NAMIBIA CRITICAL METALS SGS BATEMAN MSA THE MSA GROUP

QUBEKA MINING CONSULTANTS

CREO ENGINEERING SOLUTIONS

Knight Piésold CONSULTING

NI 43-101 TECHNICAL REPORT

ON THE

LOFDAL HEAVY RARE EARTHS PROJECT 2B-4 PRELIMINARY FEASABILITY STUDY (PFS) NAMIBIA

Prepared for:

Namibia Critical Metals Inc.

Report Date: 12 January 2026

Effective Date: 3 December 2025

Qualified Persons

Jeremy C Witley, Pr. Sci. Nat.

Joseph M. Keane, (B.S, M.S., PE Metallurgy)

Peter Christians, FAusIMM

Etienne Alain Roux, B.Eng (Chem), SME-RM

Veronique Daigle, Eng./Pr. Eng.

William van Breugel, P. Eng.

Company

The MSA Group (Pty) Ltd

SGS North America Inc.

Qubeka Mining Consultants

Creo Engineering Solutions (Pty) Ltd

Knight Piésold Consulting (Pty) Ltd

SGS Canada Inc.

SGS Bateman (Pty) Ltd

Stoneridge Office Park, Building E, Ground Floor, 8 Greenstone Place, Edenvale, 1609, South Africa

t +27(0)10 900 1990 www.za.sgs.com Company Registration Number: 1980/003077/07

Member of the SGS Group (SGS SA) Directors: L.G.B Grouiller; G. Van Aswegen; G. M Corlett; T.R.O. Abasov

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Disclaimer:

This document is issued by the SGS Bateman (Pty) Ltd. under its General Conditions of Service accessible at http://www.sgs.com/terms_and_conditions.htm. Attention is drawn to the limitation of liability, indemnification and jurisdiction issues defined therein. Any holder of this document is advised that information contained herein reflects the Company's findings at the time of its intervention only and within the limits of the Client's instructions, if any. The Company's sole responsibility is to its Client and this document does not exonerate parties to a transaction from exercising their rights and obligations under the transaction documents. Any unauthorized alteration, forgery or falsification of the content or appearance of this document is unlawful and offenders may be prosecuted.

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TABLE OF CONTENTS ... ii
LIST OF FIGURES ... xiv
LIST OF TABLES ... xix
1. Summary ... 23
1.1. Introduction ... 23
1.2. Accessibility, Climate, Local Resources, Infrastructures, and Physiography ... 23
1.3. Geology and Mineralization ... 23
1.4. Exploration Status ... 24
1.5. Mineral Processing and Metallurgical Testing ... 26
1.6. Mineral Resource Estimate ... 28
1.7. Mineral Reserve Estimate ... 30
1.8. Mining Methods ... 31
1.9. Recovery Methods ... 32
1.10. Project Infrastructure ... 32
1.11. Rare Earth Pricing ... 34
1.12. Socio-Economic and Environmental Impact ... 35
1.13. Legal and Statutory ... 36
1.14. Capital and Operating Cost Estimate ... 36
1.15. Economic Analysis ... 37
1.16. Conclusions ... 41
1.17. Recommendations ... 41
2. Introduction ... 43
2.1. Purpose of Report ... 43
2.2. Terms of Reference ... 44
2.3. Qualifications of Consultants ... 44
2.4. Report Responsibility and Qualified Persons ... 44
2.5. Site Visit ... 45
2.6. Currency, Units, Abbreviations and Definitions ... 45
2.7. Effective Date ... 47
2.8. Previous Technical Reports ... 47
3. Reliance on other Experts ... 48
4. Property Description and Location ... 49
4.1. Property Location ... 49
4.2. Property Description ... 49
4.2.1. Mining License (ML) 200 ... 49
4.2.2. General Provisions ... 53
4.2.3. Adjacent and Overlapping EPLs ... 54
5. Accessibility, Climate, Local Resources, Infrastructure and Physiography ... 55
5.1. Accessibility ... 55
5.2. Climate ... 55

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5.3. Local Resources and Infrastructure ... 56
5.4. Physiography ... 57

History ... 59
Geological Setting and Mineralization ... 62

7.1. Regional Geology ... 62
7.2. Local Geology ... 64

7.2.1. The Huab Metamorphic Complex ... 66
7.2.2. The Damara Orogen ... 66
7.2.3. Early Damaran Alkaline / Carbonatitic Intrusions ... 67

7.2.3.1. The Oas Syenite ... 67
7.2.3.2. The Lofdal Carbonatite Complex ... 68

7.2.3.2.1. Nepheline Syenite ... 68
7.2.3.2.2. Phonolite ... 68
7.2.3.2.3. Lofdal Breccias ... 69
7.2.3.2.4. Carbonatites ... 70

7.3. Structural Setting ... 72
7.4. REE Mineralization ... 74

7.4.1. Regional Setting ... 74
7.4.2. Mineralization in Area 4 ... 76
7.4.3. Mineralization in Area 2B ... 78
7.4.4. Nature of the Alteration ... 81
7.4.5. Mineralogy ... 86
7.4.6. Thorium ... 88
7.4.7. Mineralization Summary ... 89

Deposit Types ... 90

8.1. General Models for REE Mineralization in Carbonatites ... 90
8.2. Magmatic Mineralization ... 91
8.3. Hydrothermal Mineralization ... 91

Exploration ... 93

9.1. Copper – Gold Exploration: 2006 – 2008 ... 93
9.2. Regional Assessment of Rare Earth Element Potential ... 93

9.2.1. Geological and Lithogeochemical Survey ... 93
9.2.2. Remote Sensing and Regional Geophysics ... 96
9.2.3. Regional Geological Mapping ... 98

9.3. Target Exploration in Area 2B ... 99

9.3.1. Geological Mapping and Lithogeochemistry ... 99

9.3.1.1. Trenching ... 101

9.4. Target Exploration in Area 4 ... 105

9.4.1. Geological Mapping and Surface Sampling ... 105
9.4.2. Ground Geophysics ... 105
9.4.3. Trenching ... 106

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Drilling ... 110
10.1. Area 2B, 2010 and 2011 Drilling ... 110
10.2. Area 4 Mineral Resource Drilling, 2011 and 2012 ... 113
10.3. Areas 4 and 2B Mineral Resource Drilling, 2020 ... 114
10.3.1. Area 4 and 2B Diamond Drilling Procedures ... 115
10.3.2. Core Recovery ... 116
10.3.3. Collar and Downhole Surveys ... 116
10.4. Areas 4 and 2B Reverse Circulation Drilling, 2023 ... 117
10.5. Interpretation of Drilling Results ... 119
10.5.1. Area 4 ... 119
10.5.2. Area 2B ... 121
10.6. Exploration Drilling Outside the Mineral Resource Areas ... 122
10.6.1. Location and Procedures ... 122
10.6.2. Exploration Drilling Results ... 127
10.6.2.1. Area 2 ... 127
10.6.2.2. Area 4 NE Extension ... 127
10.6.2.3. Area 5 ... 128
10.6.2.4. Area 6 ... 128
10.6.2.5. Area 8 ... 128
10.6.2.6. Dolomite Hill ... 128
10.6.2.7. North Splay ... 128
Sample Preparation, Analyses and Security ... 130
11.1. Diamond Drilling Procedures ... 130
11.1.1. Drillhole Logging ... 130
11.1.2. Sample Preparation ... 131
11.1.2.1. Core Marking and Splitting ... 131
11.1.2.2. Core Sampling and Sample Dispatch ... 133
11.1.2.3. Density Measurements ... 134
11.1.2.4. Core Storage ... 135
11.2. Reverse Circulation Drilling Procedures ... 135
11.2.1. Reverse Circulation Logging ... 135
11.2.2. Chip Sampling and Sample Dispatch ... 136
11.3. Sample Analyses ... 136
11.3.1. Sample Preparation at the Laboratory ... 136
11.3.2. Sample Analyses at the Laboratory ... 136
11.4. Sample Security ... 137
11.5. Quality Assurance and Quality Control ... 137
11.5.1. 2010, 2011 and 2012 Drilling Programme ... 138
11.5.2. 2020 Drilling Programme ... 138
11.5.2.1. Blank Samples ... 138
11.5.2.2. Standards ... 140

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11.5.2.3. Pulp Duplicates ... 148
11.5.2.4. Second Laboratory Duplicate Assays ("Umpire Laboratory") ... 149
11.5.3. 2023 RC Programme ... 151
11.5.3.1. Blanks ... 151
11.5.3.2. Standards ... 152
11.5.3.3. Pulp Duplicates ... 155
11.5.3.4. Field Duplicates ... 156
11.6. Adequacy of Sample Preparation, Security and Analytical Procedures ... 158
12. Data Verification ... 159
13. Mineral Processing and Metallurgical Testing ... 161
13.1. Historical Testwork Background ... 161
13.2. Ongoing Testwork ... 161
13.2.1. Ore Sorting ... 161
13.2.2. Physical Processing ... 162
13.2.3. Hydrometallurgical Processing ... 170
13.2.4. Further Testwork ... 181
13.3. Basis of Design ... 181
13.3.1. Recommendation for the Preferred Flowsheet ... 181
13.3.2. Risks and Opportunities ... 182
14. Mineral Resource Estimates ... 183
14.1. Mineral Resource Estimation Database ... 183
14.2. Exploratory Analysis of the Raw Data ... 186
14.2.1. Validation of the Data ... 186
14.2.2. Statistics of the Raw Sample Data ... 187
14.2.2.1. Sample Lengths ... 187
14.3. Bivariate Analysis ... 188
14.4. Core Recovery ... 188
14.4.1. Diamond Drillholes ... 188
14.4.2. Reverse Circulation ... 188
14.5. Geological Modelling ... 189
14.5.1. Topography ... 189
14.5.2. Mineralised Zones ... 189
14.5.3. Oxidation/Weathering Surface ... 191
14.6. Bias Test ... 192
14.7. Statistical Analysis of the Composite Data ... 193
14.7.1. Cutting and Capping ... 196
14.8. Geostatistical Analysis ... 196
14.8.1. Semi-Variograms ... 196
14.9. Block Modelling ... 199

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14.9.1. Estimation Parameters ... 200
14.10. Validation of Estimates ... 201
14.11. Mineral Resource Classification ... 204
14.12. Mineral Resource Statement ... 206
14.13. Assessment of Reasonable Prospects for Eventual Economic Extraction (RPEEE) 211
14.14. Comparison with Previous Estimate ... 214

Mineral Reserve Estimates ... 216
15.1. Overview ... 216
15.2. Project Tenure ... 216
15.3. Mineral Resources Considered for Mining ... 217
15.4. Mineral Reserve Estimate ... 217
Mining Methods ... 218
16.1. Overview ... 218
16.2. Geotechnical Evaluation ... 218
16.3. Hydrogeological Evaluation ... 220
16.4. Open Pit Optimisation ... 222
16.4.1. Summary Optimisation Parameters ... 222
16.4.2. Geological Block Models Input to Pit Optimisation ... 223
16.4.3. Processing Plant Capacity ... 227
16.4.4. General & Administration (Fixed) Costs ... 227
16.4.5. Processing Recovery ... 227
16.4.6. Mining Costs ... 228
16.4.7. Taxes, Levies and Selling Costs ... 229
16.4.7.1. Selling Costs ... 229
16.4.7.2. Royalties ... 229
16.4.8. Processing Costs ... 230
16.4.9. Open Pit Constraints and Mining Limits ... 230
16.4.10. Mining Recovery and Dilution ... 230
16.4.11. Product Prices ... 230
16.4.12. Net Revenue ... 230

16.5. Optimisation Results ... 231
16.5.1. Pit A2B ... 232
16.5.2. Pit A4 ... 234
16.5.3. Final Pit and Interim Shells Selection ... 238
16.5.3.1. A2B Pit Selection ... 238
16.5.3.2. A4 Pit Selection ... 238

16.6. Open Pit Design ... 241
16.6.1. Pit Ramps Design Criteria ... 241
16.6.2. Pit A2B Design ... 242
16.6.3. Pit A4 Design ... 243

16.7. Dump Design ... 245

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16.8. Dewatering ... 247
16.9. Operating Hours ... 247
16.10. Mining Equipment ... 247
16.11. Mining Personnel ... 248
16.11.1.1. Owner's Mining Labour Complement ... 248
16.11.1.2. Mining Contractor Labour Complement ... 249

16.12. Production Schedule ... 250
16.12.1. Scheduler Setup ... 250
16.12.2. Schedule Control ... 251
16.12.3. Schedule Targets ... 252
16.12.4. Schedule Results ... 252
16.12.5. Haul Network and Cycle Times ... 265
16.12.6. Yearly Progress Plots ... 265

Recovery Methods ... 276
17.1. Process Design Basis ... 278
17.2. Process Design Criteria, Summary ... 279
17.3. Plant Description ... 283
17.3.1. Concentrator ... 284
17.3.1.1. Ore Reception Crushing and Sorting ... 284
17.3.1.2. Milling (Ball Milling) ... 284
17.3.1.3. Flotation ... 285
17.3.1.4. Concentrate Dewatering and Filtration ... 285
17.3.1.5. Tailings Dewatering and Transfer ... 285
17.3.2. Refinery ... 285
17.3.2.1. Sulphation Roast ... 285
17.3.2.2. REE Water Leach ... 286
17.3.2.3. Impurities Precipitation and Filtration ... 286
17.3.2.4. Uranium Ion Exchange ... 286
17.3.2.5. ADU Precipitation ... 287
17.3.2.6. Rare Earth Elements Precipitation and Thickening ... 287
17.3.2.7. Rare Earth Elements Carbonate Precipitate Drying and Packaging ... 287
17.3.3. Reagents ... 287
17.3.3.1. Sulphuric Acid ... 287
17.3.3.2. Magnesium Carbonate ... 287
17.3.3.3. Sodium Carbonate ... 288
17.3.3.4. Ammonium Hydroxide ... 288
17.3.3.5. Milk of Lime (Calcium Hydroxide) ... 288
17.3.3.6. Primary (3900z) and Secondary (3000) Collectors ... 288
17.3.3.7. Flocculant ... 289
17.3.3.8. Sodium Hydroxide ... 289
17.3.3.9. Calgon ... 289
17.3.3.10. Sodium Silicate ... 289
17.3.4. Utilities ... 289

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17.3.4.1. Plant Air and Instrument Air (Compressed Air) ... 289
17.3.4.2. Process Water ... 290
17.3.4.3. Raw Water ... 290
17.3.4.4. Potable Water ... 290
17.3.4.5. Gland Seal Water ... 290
17.3.4.6. Fire Water ... 290
17.3.4.7. Demineralized Water ... 290
17.3.4.8. Steam ... 291
17.3.4.9. Fuel ... 291

17.4. Safety and Risk Assessment ... 291
17.4.1. Background and Objectives ... 291
17.4.2. Risk Management Process ... 291
17.4.3. Risk Identification and Mitigation ... 293

17.5. Recommendations ... 297

Project Infrastructure ... 298
18.1. Summary ... 298
18.2. Bulk Water Supply ... 299
18.2.1. Water Demand ... 299
18.2.2. General ... 300
18.2.3. Boreholes ... 300
18.2.4. Transfer Pipeline ... 301
18.2.5. Pipeline Appurtenances ... 302
18.2.5.1. Inline Isolating Valve Installations ... 302
18.2.5.2. Scour Installations ... 302
18.2.5.3. Air Valve Installations ... 302
18.2.5.4. Pipeline Markers ... 302

18.2.6. Water Storage along the Pipeline Route ... 302
18.2.6.1. Borehole Collector and Pressure Break Reservoirs ... 302

18.2.7. Motor Control Centres ... 303
18.2.7.1. General ... 303
18.2.7.2. Borehole Pump Control ... 303
18.2.7.3. PLC and SCADA ... 304

18.2.8. Electrical Power Supply ... 304
18.2.9. Surge / Lightning Protection, Earthing and Bonding ... 304
18.2.10. Fire Protection ... 304
18.2.11. Pump Station Buildings / Enclosures / Structures ... 305
18.2.12. Raw Water Storage ... 305
18.2.12.1. General ... 305
18.2.12.2. Fire Water Storage ... 305

18.2.13. Domestic Water Supply ... 306
18.2.13.1. General ... 306
18.2.13.2. Water Treatment Plant ... 306
18.2.13.3. Potable Water Booster Pump ... 306

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18.2.13.4. Potable Water Storage ... 307
18.2.13.5. External Potable Water Reticulation ... 307
18.2.13.6. Internal Potable Water Reticulation ... 307

18.3. Bulk Power Supply ... 307
18.3.1. Electrical Load ... 307
18.3.2. Power Generation ... 307
18.3.3. Transmission Line ... 308
18.3.4. 132/11 kV Lofdal Substation ... 310
18.3.5. Electrical Works ... 311
18.3.6. IPP Supply ... 311
18.3.7. Solar PV ... 311
18.3.8. Diesel Generation ... 312
18.3.9. Battery Energy Storage System (BESS) ... 313

18.4. Electrical Distribution ... 313
18.5. Site-Wide Communications ... 313
18.6. Access Roads ... 314
18.7. Plant Buildings ... 315
18.8. Transportation and Site Vehicles ... 316
18.8.1. Site LDVs ... 316
18.8.2. Personnel Transportation ... 317
18.8.3. Material Transport to and from Site ... 318

18.9. Plant Fencing ... 319
18.10. Tailings Storage Facility ... 319
18.10.1. TSF Concept Design ... 320
18.10.2. Storage Capacity Assessment ... 320
18.10.3. Foundation Characteristics ... 321
18.10.4. Tailings Material Testing ... 322
18.10.5. Design Criteria ... 323
18.10.6. TSF Components and Geometry ... 324
18.10.6.1. Key Quantities ... 325

18.10.7. Dam Safety Classification ... 326
18.10.7.1. Dam Breach Analysis ... 326
18.10.7.2. Dam Safety Classification – Consequence of Failure ... 327

18.10.8. Preliminary Seepage and Stability Analysis ... 327
18.10.9. TSF Development and Operational Philosophy ... 329
18.10.10. TSF Water Management ... 330
18.10.10.1. Climate ... 330
18.10.10.2. Rainfall and Evaporation ... 330
18.10.10.3. Extreme Rainfall Estimation ... 330
18.10.10.4. TSF Stormwater Management ... 331
18.10.10.5. Water Balance ... 331

18.11. Plant Sewerage Treatment & Distribution ... 334
18.12. Design Parameters ... 335

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Market Studies and Contracts ... 336
19.1. Introduction ... 336
19.2. Overview of Rare Earth Elements ... 336
19.3. Demand for Magnet Rare Earths and Yttrium ... 337
19.3.1. Magnet Demand Trends ... 337
19.3.2. Yttrium Demand and Strategic Importance ... 337
19.3.3. Global REE Demand by 2040 ... 338
19.4. Supply, Geopolitics and Yttrium Export Controls ... 340
19.4.1. Chinese Dominance and Export Controls ... 340
19.4.2. Price Impacts and “Scramble for Yttrium” ... 340
19.5. Recent Market Pricing and Peer Price Decks ... 341
19.5.1. Emerging Ex-China Price Divergence ... 341
19.5.2. Drivers of Price Divergence ... 341
19.5.3. Divergence Price Trends ... 342
19.5.4. Implications for Current and Forecast Pricing for Lofdal ... 342
19.5.5. Indicative Long-term View on Yttrium Prices ... 343
19.6. Marketing Strategy ... 343
19.7. Existing Contracts ... 344
19.8. Indicative Price Framework and Basket Value ... 344
19.9. Conclusions ... 345
Environmental Studies, Permitting and Social or Community Impact ... 347
20.1. Project Background ... 347
20.2. Motivation for the Project ... 350
20.3. Project Description Overview ... 350
20.4. Project Alternatives ... 354
20.5. EIA and Public Consultation Process ... 354
20.6. Profile of the Receiving Environment ... 355
20.6.1. Biophysical Environment ... 355
20.6.2. Cultural Heritage ... 357
20.6.3. Socio-Economic ... 358
20.7. Specialist Studies ... 360
20.8. Quantification of Impacts ... 360
20.9. Mitigation ... 366
20.10. Environmental Statement and Conclusion ... 366
Capital and Operating Costs ... 367
21.1. Basis of Estimate ... 367
21.2. Project Capital Summary ... 367
21.2.1. Mining ... 368
21.2.2. Process Capital ... 369
21.2.2.1. Estimating Approach ... 369
21.2.2.2. Qualifications, Assumptions, Exclusions and Exceptions ... 369

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21.2.2.3. Estimating Criteria ... 369
21.2.2.3.1. Capital Cost Estimate ... 369
21.2.2.4. Estimating Accuracy ... 370
21.2.2.5. Base Date ... 370
21.2.2.6. Scope Definition ... 370
21.2.2.7. Pricing Basis ... 370
21.2.2.8. Presentation of Capital Cost ... 370
21.2.2.9. Capital Estimate Structure ... 371
21.2.2.10. Direct Field Costs (DFC) ... 371
21.2.2.10.1. Earthworks ... 371
21.2.2.10.2. Civil Works ... 371
21.2.2.10.3. Buildings and Infrastructure ... 371
21.2.2.10.4. Structural Steel ... 371
21.2.2.10.5. Platework and Liners ... 371
21.2.2.10.6. Mechanical Equipment ... 372
21.2.2.10.7. Piping and Valves ... 372
21.2.2.10.8. Electrical ... 372
21.2.2.10.9. Control and Instrumentation ... 372
21.2.2.10.10. Preliminary and General ... 373
21.2.2.10.11. Allowances ... 373
21.2.2.10.12. Transportation and Logistics ... 373
21.2.2.10.13. Spares ... 373
21.2.2.10.14. First Fills (Oils, Lubricants) ... 373
21.2.2.10.15. Vendor Assistance ... 373
21.2.2.11. Indirect Field Costs (IFC) ... 374
21.2.2.11.1. Engineering, Design and Project Management ... 374
21.2.2.11.2. Bonds, Guarantees etc. ... 374
21.2.2.11.3. Project Insurance ... 374
21.2.2.11.4. Project Contingency ... 375
21.2.2.12. Owner's Cost ... 375
21.2.2.13. Exclusions ... 375
21.2.3. Tailings Capital Costs ... 378
21.2.4. Non-Process Infrastructure ... 378
21.3. Project Operating Costs ... 380
21.3.1. Process Plant Operating Cost Estimate ... 380
21.3.1.1. Scope ... 381
21.3.1.2. Accuracy ... 382
21.3.1.3. Exclusions ... 382
21.3.2. Fixed Costs ... 382
21.3.2.1. Labour ... 382
21.3.2.2. Maintenance ... 383
21.3.3. Variable Costs ... 383
21.3.3.1. Power ... 383
21.3.3.2. Reagents and Consumables ... 383
21.3.3.3. Fuel ... 384
21.3.3.4. Power and Water Cost ... 385

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21.3.3.5. Site Closure...385
21.4. Mine Operating Costs...385
21.5. General and Administration...388
22. Economic Analysis...389
22.1. Introduction...389
22.2. Basis of Economic Analysis...390
22.2.1. Project Contingency...390
22.2.2. Rare Earth Pricing Models...391
22.2.3. Clarification and Assumptions...391
22.2.4. Analysis Period...391
22.2.5. Operating Costs...391
22.2.6. Capital Costs...391
22.2.7. Funding...392
22.2.8. After Tax Free Cash Flow...392
22.2.9. Net Present Value...392
22.2.10. After Tax Internal Rate of Return...392
22.2.11. Payback Period...392
22.3. Base Case Financial Model and Sensitivities...392
22.3.1. Base Case Project Economics...392
22.3.2. Base Case Sensitivity Analysis...396
22.4. Divergent Pricing Financial Model and Sensitivities...398
22.4.1. Divergent Pricing Project Economics...398
22.4.2. Divergent Pricing Sensitivity Analysis...401
22.4.3. Recovery Sensitivities...403
22.4.4. Discussion on Sensitivities...403
22.4.5. Conclusions and Recommendations...403
23. Adjacent Properties...404
24. Other Relevant Data and Information...405
25. Interpretation and Conclusions...406
25.1. Mineral Resource Estimate...406
25.2. Capital and Operating Cost...407
25.3. Economic Analysis...408
25.3.1. Sensitivity Analysis...408
25.3.1.1. After-Tax Price, Cost & Exchange Rate Sensitivities...409
25.3.2. Overall Economic Interpretation...410
25.4. Opportunities...410
25.5. Risks...410
26. Recommendations...412
26.1. Metallurgical Testwork...412
26.2. Mining...412
26.3. Mineral Processing...412

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26.4. Infrastructure... 413
26.4.1. Water Supply... 413
26.4.2. Electrical Supply... 413
26.4.3. Mine Access Road... 413
26.4.4. Tailings Storage Facility... 413
26.5. REE Pricing... 413
26.6. Economic Analysis... 414
26.7. Overall... 414
27. References... 415
28. Date and Signature Page... 418
29. Certificates of Qualified Persons... 420

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LIST OF FIGURES

Figure 1 Lofdal Area 4 pit as of March 2025 ... 25
Figure 2 Simplified hydrometallurgical flowsheet developed during PFS ... 28
Figure 3 Layout of planned mining and processing infrastructure at Lofdal ... 33
Figure 4 Illustrative cash flow profile for Base Case and Divergent Case; year-on-year variability occurs in practice ... 39
Figure 5 Sensitivities for Base Case and Divergent Case ... 40
Figure 6 Process Recovery Sensitivities for Base Case and Divergent Case ... 41
Figure 7 Location of the Lofdal Property (red square NW of Khorixas) ... 49
Figure 8 Location of ML 200 and the boundary of EPL3400 at the time it was relinquished showing current boundaries, roads, and the location of the Hoppe Mineral Claims and Mining License Applications ... 50
Figure 9 Location of Mining License 200 (blue line). Corner numbers in red are same as given in Table 11 ... 52
Figure 10 Extent of Mining License 200 and the Former EPL 3400 in relation to the Mineral Resources of Area 4 and Area 2B ... 53
Figure 11 Location and road access to the Lofdal project area ... 55
Figure 12 Facilities in Khorixas for equipment storage, core logging and storage ... 56
Figure 13 Core processing facilities at the Lofdal Field Camp ... 57
Figure 14 Physiography of the project area showing typical low rolling hills and sparse vegetation ... 58
Figure 15 Adit at the former Lofdal Copper Mine with copper staining around the portal ... 59
Figure 16 Pit sampling of a carbonatite dyke from Rouna’s exploration ... 60
Figure 17 Cratons and orogenic belts in southern Africa ... 62
Figure 18 General geology of Namibia ... 63
Figure 19 General geology in the area of the Welwitschia Inlier ... 64
Figure 20 Detailed geology of the area of the Lofdal Carbonatite complex ... 65
Figure 21 General topography and outcrop appearance of the Huab metamorphic Complex ... 66
Figure 22 Coarse grained Oas Syenite (alkali feldspar, amphibole and mica) ... 67
Figure 23 Examples of nepheline syenite and phonolite dyke ... 69
Figure 24 Examples of Lofdal breccias ... 70
Figure 25 Examples of carbonatite from the Main and Emanya intrusions ... 71
Figure 26 Examples of brown and red to yellow carbonatite dykes ... 72
Figure 27 Structural elements of the Lofdal area, interpreted from Landsat and hyperspectral data ... 73
Figure 28 Distribution of lithogeochemical grab samples in the Lofdal area ... 74
Figure 29 Lithogeochemical grab samples plotted on the basis of (HREE+Y)/(TREE+Y) ... 75
Figure 30 Geology of Area 4 with Dysprosium (Dy) grade in surface grab samples ... 77
Figure 31 (HREE+Y)/(TREE+Y)% in surface grab samples in Area 4 ... 78
Figure 32 Geology of Area 2B with dysprosium (Dy) grade in surface grab samples ... 79
Figure 33 (HREE+Y)/(TREE+Y) % in surface grab samples in Area 2 ... 80
Figure 34 Colour anomaly in drillhole core associated with the Main zone alteration in Area 4 .82
Figure 35 Schematic illustration of geochemical and radiometric anomalies associated with the Area 4 alteration zone in drillhole NLOFDH4047 ... 82
Figure 36 Schematic cross section of the upper part of the Area 4 Main alteration zone from drillhole data ... 83
Figure 37 Typical alteration and mineralization in Area 2B ... 84
Figure 38 Examples of Area 4 alteration in drillhole core ... 85

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Figure 39 High grade mineralization in Area 4 alteration in drillhole core...85
Figure 40 Albitite with aggregates of xenotime and zircon...86
Figure 41 Backscatter images of Area 4 mineralization...87
Figure 42 THREE versus Th in trench and drillhole core samples...88
Figure 43 General cross-sectional model for an alkali silicate-carbonate intrusive complex (after Le Bas, 1987)...90
Figure 44 Distribution of regional lithogeochemical samples in the Lofdal area...94
Figure 45 Priority exploration areas defined by Dy in surface lithogeochemistry samples...95
Figure 46 Priority exploration areas defined by HREE/TREE ratio in surface lithogeochemistry samples...96
Figure 47 Trenches on the Area 2B Zone (heavy brown lines) – trenches listed in Table 14 are labelled...101
Figure 48 A – digging trenches on Area 2B with a backhoe, B – cleaning trenches in preparation for sampling and mapping...102
Figure 49 Area 4 grid showing ground geophysical coverage...105
Figure 50 Location of Trenches in Area 4...106
Figure 51 Examples of trenches in Area 4...107
Figure 52 Location of drillhole collars in Area 2B...112
Figure 53 Plan showing collars of the 101 drillholes drilled in 2011 and 2012 in Area 4...114
Figure 54 Plan showing drillhole collars in Area 4...115
Figure 55 Drillers’ metre marks, measured metre marks and orientation lines on uncut core...116
Figure 56 Concrete plinth over capped diamond drillhole L4D017 (left) and closed off RC hole L4R0215 (right)...117
Figure 57 RC Drilling at Area 4 (hole number L4R0218)...118
Figure 58 Plan showing Area 4 drillhole collars...119
Figure 59 Plan showing Area 2B collars...119
Figure 60 Example of a drilling section through the Main Zone mineralization at Area 4...121
Figure 61 Example of drilling section through the Area 2B mineralized zone...122
Figure 62 Location of exploration (non-resource) drillholes (white squares)...123
Figure 63 Illustration of the alpha, beta, and gamma angles in core...130
Figure 64 Examples of drillhole core marking...132
Figure 65 Core cutting device...133
Figure 66 Apparatus for measuring density (SG) by the Archimedes principle...134
Figure 67 Core storage in the Khorixas warehouse and in containers in the warehouse yard...135
Figure 68 Blank analyses for selected REE from Area 2B...139
Figure 69 Blank analyses for selected REE from Area 4...140
Figure 70 Analyses of the CRM AMIS0185 for selected REE in Area 2B...141
Figure 71 Analyses of the CRM AMIS0185 for selected REE in Area 4...142
Figure 72 Analyses of the Standard STD4 for selected REE in Area 2B...143
Figure 73 Analyses of the Standard STD4 for selected REE in Area 4...143
Figure 74 Analyses of the Standard STD5 for selected REE in Area 4...144
Figure 75 Analyses of the Standard STD6 for selected REE in Area 2B...144
Figure 76 Analyses of the Standard STD6 for selected REE in Area 4...145
Figure 77 Analyses of laboratory duplicates for selected REE from Area 2B...148
Figure 78 Analyses of laboratory duplicates for selected REE from Area 4...149
Figure 79 Scatterplot of duplicate pair data (Actlabs and the umpire lab (ALS)) for selected REES...150
Figure 80 2023 RC programme blank chart for dysprosium...151
Figure 81 2023 RC programme blank chart for cerium...152

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Figure 82 Dysprosium standard control charts...153
Figure 83 Scatterplot of TREO assays in 285 pulp duplicates...155
Figure 84 Scatterplot of Dy₂O₃ assays in 285 pulp duplicates...156
Figure 85 Scatterplot of TREO assays for 35 field duplicates...157
Figure 86 Scatterplot of TREO assays for 35 field duplicates...157
Figure 87 Pilot Plant Flowsheet...163
Figure 88 Flotation Variability Testing Additional Results...166
Figure 89 Mineralogy - Liberation of Variability Flotation Cleaner Tailings...167
Figure 90 Flotation Variability Testing Finer Grinds...168
Figure 91 Further Flotation Results on ROM Bulk Sample...169
Figure 92 Selected Best Flotation Results of Sorted Feed Samples...170
Figure 93 Acid bake performance at different temperatures...171
Figure 94 Acid bake performance using different acid addition at 600°C...172
Figure 95 Comparison between single and two-stage static acid bake process...172
Figure 96 Effect of acid dosage on two-stage acid bake process...173
Figure 97 Comparison between single and two-stage dynamic acid bake process...174
Figure 98 Initial block flow diagram of hydrometallurgical flowsheet before the PFS test work...174
Figure 99 Metal precipitation as a function of pH and temperature...176
Figure 100 Metal precipitation behaviour of WL slurry and WL filtrate...176
Figure 101 Secondary neutralisation of pH...177
Figure 102 Dysprosium precipitation and tenors as a function of pH...178
Figure 103 Rare earth precipitation pilot plant flowsheet...178
Figure 104 Simplified flowsheet developed during further downstream hydrometallurgical testing...180
Figure 105 Recommended flowsheet...182
Figure 106 Collar positions by campaign for Area 4...185
Figure 107 Collar positions by campaign for Area 2B...185
Figure 108 Cumulative frequency distribution comparison between pre-2020 downhole probe and 2020 Archimedes densities...187
Figure 109 Histogram of DD sample lengths for Area 4 and Area 2B...187
Figure 110 Scatter plot of sample Tb₂O₃, Dy₂O₃ and Ho₂O₃ for Area 4...188
Figure 111 Histogram of RC sample weights for Area 4 (left) and Area 2B (right)...189
Figure 112 Cross-section illustrating modelled mineralised zones for Area 4...190
Figure 113 Cross-section illustrating modelled mineralised domains for Area 2B...191
Figure 114 Cross-section illustrating modelled RQD weathering surface for Area 4...192
Figure 115 Histograms of TREO, LREO, HREO and Dy₂O₃ ppm for MZONE 1 in Area 4...194
Figure 116 Histograms of TREO, LREO, HREO and Dy₂O₃ ppm for MZONE 1 in Area 2B...195
Figure 117 Semi-variogram models for HREO in MZONE 1 - Area 4...199
Figure 118 Swath Plot Validation for Dy₂O₃ – Area 4 MZONE 1...202
Figure 119 Area 4 block model cross-section coloured on Dy₂O₃...203
Figure 120 Area 2B block model cross-section coloured on Dy₂O₃...204
Figure 121 Mineral Resource classification for Area 4...205
Figure 122 Mineral Resource classification for Area 2B...206
Figure 123 Area 4 – plan showing block model relative to pit shell extents...212
Figure 124 Area 4 section looking northeast showing block model relative to pit shell extents and topography...213
Figure 125 Area 2 – plan showing block model relative to pit shell extents...213

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Figure 126 Area 2 section looking northeast showing block model relative to pit shell extents and topography...214
Figure 127 Location of ML 200 and EPL3400 at relinquishment...216
Figure 128 Historical 2023 and 2025 drillhole locations (SRK)...219
Figure 129 Geotechnical Design Sectors...219
Figure 130 Groundwater monitoring boreholes at the project area (SLR)...221
Figure 131 GWL of quarterly monitored boreholes at Lofdal (SLR)...221
Figure 132 Grade-Tonnage Curves for Regularised Models vs. Sub-blocked Models...226
Figure 133 Process Plant Flowsheet...227
Figure 134 A4 Project Evaluator Task Setup...232
Figure 135 A2B Best/Worst Case Analysis...233
Figure 136 A2B Pit Shell Strip Ratio and Grade...234
Figure 137 Pit A4 Best-Worst Case Analysis...235
Figure 138 A4 Pit Shell Strip Ratio and Grade...236
Figure 139 Specified Case Analysis...237
Figure 140 Final pit shells selected for A2B and A4 pits...239
Figure 141 A4 Interim Pushbacks...240
Figure 142 Pit Ramp Design Criteria...241
Figure 143 Pit A2B Final Pit Design...242
Figure 144 Pit A4 Interim Pushbacks and Final Pit Designs...243
Figure 145 Dump Design Parameters...245
Figure 146 Waste Rock Dumps & Stockpiles...246
Figure 147 Scheduler (MPSO) Process Flow...251
Figure 148 Process Throughput and Recovery Ramp-up...252
Figure 149 Pre-strip and First 3-Years Quarterly Pit Production Schedule...254
Figure 150 LOM Annual Pit Production Schedule...256
Figure 151 ROM Mined by Resource Class...259
Figure 152 Schedule Haulage Setup...265
Figure 153 Year-00 (Pre-Strip) Pit A4 Progress...266
Figure 154 Year-01 Pit A4 Progress...266
Figure 155 Year-02 Pit A4 Progress...267
Figure 156 Year-03 Pit A4 Progress...267
Figure 157 Year-04 Pit A4 Progress...268
Figure 158 Year-05 Pit A4 Progress...268
Figure 159 Year-05 Pit A2B Progress...269
Figure 160 Year-06 Pit A4 Progress...269
Figure 161 Year-06 Pit A2B Progress...270
Figure 162 Year-07 Pit A4 Progress...270
Figure 163 Year-07 Pit A2B Progress...271
Figure 164 Year-08 Pit A4 Progress...271
Figure 165 Year-08 Pit A2B Progress...272
Figure 166 Year-09 Pit A4 Progress...272
Figure 167 Year-09 Pit A2B Progress...273
Figure 168 Year-10 Pit A4 Progress...273
Figure 169 Year-10 Pit A2B Progress...274
Figure 170 Year-11 Pit A4 Progress...274
Figure 171 Year-12 Pit A4 Progress...275
Figure 172 Year-13 Pit A4 Progress...275

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Figure 173 Block Flow Diagram Processing Plant ... 277
Figure 174 Lofdal Project Site Location ... 298
Figure 175 Lofdal Project Site Layout ... 299
Figure 176 Assumed Borehole Locations and Pipeline Route ... 300
Figure 177 EPANET Model ... 301
Figure 178 Bulk Power Infrastructure ... 308
Figure 179 Transmission Line Route ... 308
Figure 180 Line Route Elevation ... 309
Figure 181 Proposed Line Route and Bend Points ... 310
Figure 182 132/11kV In-take Substation Single Line Diagram (incorporating future second transformer) ... 311
Figure 183 Access route via the D2625 from the North ... 314
Figure 184 Access route via the D2625 from the East. ... 315
Figure 185 Preliminary buildings layout plan ... 316
Figure 186 Preferred route for inbound and outbound material. ... 319
Figure 187 Tailings Design Layout ... 320
Figure 188 TSF Filling Curve and Estimate of Starter Wall Requirement ... 321
Figure 189 Inundation zone for rainy day scenario ... 326
Figure 190 Cross Section of the starter wall ... 327
Figure 191 Cross Section and Material of the Final Landform (tailings at 1005 mamsl) ... 328
Figure 192 Simplified Inflows and Outflows Schematic ... 331
Figure 193 Inflow, outflow and decant average monthly volumes ... 332
Figure 194 Inflow, outflow and decant wet monthly volumes ... 332
Figure 195 Inflow, outflow and decant dry monthly volumes ... 333
Figure 196 Inflow, outflow and decant summary monthly volumes ... 333
Figure 197 Global REO demand by application (in kt REO, source CRU) ... 339
Figure 198 Magnet and non-magnet demand (kt REO, source CRU) ... 340
Figure 199 Regional Setting ... 348
Figure 200 Local Setting ... 349
Figure 201 Site layout (Approved 2016 layout and amended (current) layout) ... 353
Figure 202 Opex Distribution ... 381
Figure 203 Base Case After Tax Cash Flows ... 393
Figure 204 Base Case After Tax Sensitivity ... 397
Figure 205 Divergent Pricing After Tax Cash Flows ... 398
Figure 206 Divergent Pricing Sensitivities ... 402
Figure 207 Process Recovery Sensitivities ... 403

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LIST OF TABLES

Table 1 Summary of drilling conducted at the Lofdal Project 25
Table 2 Area 4, Measured, Indicated and Inferred Mineral Resource Estimates above 0.1% TREO cut-off grade – 05 April 2024 29
Table 3 Area 2B, Indicated and Inferred Mineral Resource Estimates above 0.1% TREO cut-off grade – 05 April 2024 30
Table 4 Lofdal Mineral Reserves as of 01 December 2025 30
Table 5 Summary Pit Optimisation Parameters 31
Table 6 Capital Costs Summary of Lofdal PFS "Lofdal 2B-4" 37
Table 7 Operating Costs Summary of the Lofdal PFS "Lofdal 2B-4" 37
Table 8 Royalties and Separation Costs of Lofdal PFS "Lofdal 2B-4" 37
Table 9: Details of Site Visits and Responsibilities of the Qualified Persons 45
Table 10 List of Abbreviations 46
Table 11 Coordinates of Mining License 200 53
Table 12 Summary of remote sensing and regional geophysical surveys and interpretations 96
Table 13 Analyses of the five highest-grade surface samples in Area 2B 100
Table 14 Locational information for trenches on the 2B Zone – WGS84; UTM Zone 33S 102
Table 15 Assays from best trench intersections -NLOFTR001, NLOFTR005 and NLOFTR006, Area 2B 104
Table 16 Summary of geophysical surveys in Area 4 106
Table 17 Locational information for trenches in Area 4. WGS84, UTM Zone 33S 107
Table 18 Representative analyses from trench samples, Area 4 109
Table 19 Summary of drilling procedures for the 2010 drilling campaign in Area 2B 111
Table 20 Infill samples from 2010-2011 drillholes sampled and assayed during 2020 112
Table 21 Summary of drilling procedures for the 2011-2012 drilling campaign in Area 4 113
Table 22 Location and orientation information for exploration drillholes on the Lofdal ML. WGS84 UTM 33S 123
Table 23 Analyses of typical significant altered/mineralized intersections in exploration drillholes 129
Table 24 Number of blank failures (>10 times LDL*) 140
Table 25 Statistics for the reference materials used in the 2020 drilling program 146
Table 26 Failure rate (outside ± 3 SD) for standards assayed by Actlabs during the 2020 drilling campaign 147
Table 27 Resolution of anomalous CRM analyses 148
Table 28 Percentage of assays within mean absolute difference of 10% and 20% (above 10x LDL) – Actlabs duplicate versus original 149
Table 29 Mean and Variance of original and duplicate data – Actlabs versus ALS 150
Table 30 Failure rate (outside ± 3 SD) for standard reference material assayed by Actlabs during the 2023 RC drilling campaign 154
Table 31 Summary of lab duplicate repeatability 155
Table 32 Summary of field duplicate repeatability 156
Table 33 Pilot Plant Testing Results Summary 164
Table 34 Best Flotation Results from Variability Testing 166
Table 35 Flotation Variability - Summary of Finer Grind Flotation 168
Table 36 Selected Best Flotation Test Results on Sorted Feed Samples 169
Table 37 Summary of secondary neutralisation tests 177
Table 38 Summary of the uranium IX results 177
Table 39 Composition of the U IX barren solution 178

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Table 40 Composition of PRP and SRP products generated in Pilot Plant 179
Table 41 Summary of selected recoveries 182
Table 42 Summary of Lofdal drilling campaigns 184
Table 43 Diamond core recovery in percent per depth interval below surface 188
Table 44 Summary statistics of RC sample weights 189
Table 45 Bias test on Dy₂O₃ for Area 2B and Area 4 193
Table 46 Individual REO proportions for Area 4 and Area 2B 196
Table 47 Semi-variogram Parameters for Area 4 198
Table 48 Semi-variogram Parameters for Area 2B 198
Table 49 Block model origins Area 4 and Area 2B 200
Table 50 Search Parameters for Area 4 200
Table 51 Search Parameters for Area 2B 200
Table 52 Area 4, Measured, Indicated and Inferred Mineral Resource Estimates above 0.1% TREO cut-off grade – 5 April 2024 207
Table 53 Area 2B, Indicated and Inferred Mineral Resource Estimates above 0.1% TREO cut-off grade – 5 April 2024 207
Table 54 Area 4, Measured and Indicated Mineral Resource grade-tonnage table – 5 April 2024 208
Table 55 Area 4, Inferred Mineral Resources grade-tonnage table – 5 April 2024 208
Table 56 Area 2B, Indicated Resources grade-tonnage table – 5 April 2024 209
Table 57 Area 2B, Inferred Resources grade-tonnage table – 5 April 2024 209
Table 58 Area 4, Individual REO Measured, Indicated and Inferred Mineral Resources above 0.1% TREO cut-off grade – 5 April 2024 210
Table 59 Area 2B, Individual REO Measured, Indicated and Inferred Mineral Resources above 0.1% TREO grade – 5 April 2024 210
Table 60 Distribution of TREO in Concentrate 211
Table 61 Area 4 – 12 May 2021 Mineral Resource Estimate compared with 5 April 2024 Mineral Resource Estimate 215
Table 62 Area 2B – 12 May 2021 Mineral Resource Estimate compared with 5 April 2024 Mineral Resource Estimate 215
Table 63 Lofdal Mineral Reserves as of 01 December 2025 217
Table 64 A2B Domain Slope Design Recommendations 220
Table 65 A4 Domain Slope Design Recommendations 220
Table 66 A2B Domain Slope Design inc. Ramps 220
Table 67 A4 Domain Slope Design inc. Ramps 220
Table 68 Pit Optimisation Parameters Summary 222
Table 69 Area 2B a2bmod_20240215 Block Model 223
Table 70 Area 4 a2bmod_20240215 Block Model 223
Table 71 M+I Modifying Factors for Regularised Models 223
Table 72 Area 2B Sub-blocked Model vs. Regularised Models 224
Table 73 Area 4 Sub-blocked Model vs. Regularised Models 225
Table 74 General and Administration (Fixed) Costs 227
Table 75 Net-of-ROM Process Recovery 228
Table 76 Mining Costs 229
Table 77 Selling Cost 229
Table 78 Royalties, Taxes and Levies 229
Table 79 Processing Costs 230
Table 80 Rare Earth Product Prices 230
Table 81 Pit Optimisation Material Types 231

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Table 82 A2B Best-Worst Case Analysis ... 232
Table 83 Pit A4 Best-Worst Case Analysis ... 234
Table 84 Specified Case Analysis ... 237
Table 85 A2B Pit Shell RF=0,92 Pit Inventory ... 238
Table 86 A4 Pit Shell RF=0,98 Pit Inventory ... 239
Table 87 A4 Interim Pushbacks Summary ... 240
Table 88 A2B Pit Design vs Optimum Pit Shell ... 242
Table 89 A4 Pit Design vs Optimum Pit Shell ... 244
Table 90 Pit A4 Design Inventory Summary ... 244
Table 91 Waste Dumps and Stockpiles Capacity ... 246
Table 92 Weather Consideration ... 247
Table 93 Condensed Mining Equipment List ... 248
Table 94 Owner's Mining Labour Complement ... 248
Table 95 Mining Contractor Management and Admin Labour Complement ... 249
Table 96 Mining Contractor Production Staff Labour Complement ... 249
Table 97 Mining Contractor Maintenance Staff Labour Complement ... 250
Table 98 Mining Contractor Total Labour Complement ... 250
Table 99 Vertical Advance Rates Achieved ... 253
Table 100 Summary of the Results Tables and Figures ... 253
Table 101 Pre-Strip and First 3-Years Quarterly Pit Production Schedule ... 254
Table 102 LOM Annual Pit Production Schedule ... 255
Table 103 Pre-Strip and First 3-Years Quarterly Ore Mined by Measured and Indicated Resource Class ... 257
Table 104 LOM Annual Ore Mined by Measured and Indicated Resource Class ... 258
Table 105 LOM Ex-Pit ROM to Destinations ... 260
Table 106 LOM Stockpile Reclaim to Mill/Flotation ... 261
Table 107 LOM Marginal Stockpiled ... 261
Table 108 LOM XRT Reject to A4N Waste Rock Dump ... 262
Table 109 LOM Waste to Waste Rock Dumps ... 263
Table 110 LOM Annual Mill/Flotation Feed ... 264
Table 111 Area Codes According to Project WBS ... 278
Table 112 Major Design Parameters ... 279
Table 113 Project Risk Register ... 294
Table 114 LDV requirement on site ... 316
Table 115 Proposed models for site LDV's. ... 317
Table 116 Passenger transport to and from site. ... 317
Table 117 Inbound and outbound transport of reagents, spares, and concentrate. ... 318
Table 118 Design Criteria-Summary ... 323
Table 119 Seepage/Stability Analysis – Assumed Geotechnical Parameters ... 328
Table 120 Slope Stability Analysis Results ... 329
Table 121 Mean Monthly Rainfall and Evaporation ... 330
Table 122 24-Hour Duration Extreme Rainfall Depths Estimates ... 331
Table 123 Design Criteria Basis ... 335
Table 124 Rare Earth Applications and End-Uses fall into one of Eight End-Use Categories .. 337
Table 125 Project Data Summary that provides perspective on the scale of and amendments made to the Project ... 351
Table 126 Summary of the Profile of the Biophysical Baseline Environment ... 355
Table 127 Summary of the Profile of the Cultural Heritage Baseline Environment ... 357

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Table 128 Summary of the Profile of the Socio-economic Baseline Character 358
Table 129 ES 1-5: Summary of the Findings of the EIA 361
Table 130 Base Criteria for Project CAPEX and OPEX 367
Table 131 Capital Summary (USD) 368
Table 132 Mining Capital Cost Breakdown (USD) 369
Table 133 Capital Estimate Summary (USD) 377
Table 134 Tailings Capital Costs (USD) 378
Table 135 Non-Process Infrastructure Capital Costs (USD) 379
Table 136 Operating Cost Summary (USD) 380
Table 137 Operating Cost Estimate Nameplate Capacity 381
Table 138 Process Labour Requirements 383
Table 139 Reagent Costing 384
Table 140 Mill Site Process Plant Energy and Water Cost 385
Table 141 Mine Operating Cost Summary 386
Table 142 Annual General and Administration Costs 388
Table 143 Administration Labour 388
Table 144 Summary of Financial Analysis 389
Table 145 Project Contingency 390
Table 146 Rare Earth Oxide Pricing 391
Table 147 Base Case Financial Model 394
Table 148 Capital Cost Sensitivity 396
Table 149 Operating Cost Sensitivity 396
Table 150 Exchange Rate Sensitivity 396
Table 151 Base Case Metal Price Sensitivity 397
Table 152 Recovery Sensitivity 397
Table 153 Divergent Pricing Financial Model 399
Table 154 Capital Cost Sensitivity 401
Table 155 Operating Cost Sensitivity 401
Table 156 Exchange Rate Sensitivity 401
Table 157 Divergent Pricing Metal Price Sensitivity 402
Table 158 Process Recovery Sensitivity 402
Table 159 Area 4 Mineral Resources above a 0.1% TREO cut-off grade – 5 April 2024 406
Table 160 Area 2B Mineral Resources above a 0.1% TREO cut-off grade – 5 April 2024 407
Table 161 Capital Costs Summary of Lofdal PFS "Lofdal 2B-4" 407
Table 162 Operating Costs Summary of the Lofdal PFS "Lofdal 2B-4" 408
Table 163 Royalties and Separation Costs of Lofdal PFS "Lofdal 2B-4" 408
Table 164 Project Risks 410

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1. Summary

1.1. Introduction

The subject of this technical report is the Lofdal Heavy Rare Earth Project “Lofdal 2B-4” (“Lofdal”). The Lofdal property is covered by Mining License (ML) 200, located approximately 30 km NW of the town of Khorixas in the Kunene Region of north-western Namibia. The ML is held by Namibia Rare Earths (Pty) Ltd. (NRE (Pty)), which is 95% owned subsidiary of Namibia Critical Metals Inc. (NMI) and 5% owned by Philco 196 (Pty) Ltd, a company incorporated in Namibia as part of the mandated ownership of Historically Disadvantaged Namibians. ML 200 was granted on 15 July 2021 and is valid until 10 May 2046.

Namibia Critical Metals (NMI or the Company) is a Canadian company listed on the TSX Venture Exchange and OTCQB Market which is developing the Lofdal Heavy Rare Earth Project within the Republic of Namibia. The Company’s registered corporate office is Suite 802, Sun Tower, 1550 Bedford Highway, Halifax, Nova Scotia, NS B4A 1E6 Canada.

Lofdal is developed under a Joint Venture Agreement with Japan Organization for Metals and Energy Security (JOGMEC).

This Preliminary Feasibility Study technical report (PFS) for the Lofdal Project “2B-4” is submitted herewith as an independent qualified person’s (QP) review and according to the National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

1.2. Accessibility, Climate, Local Resources, Infrastructures, and Physiography

The Project site is located on the farm Lofdal, in the Kunene region of Central Northwest Namibia. The coordinates of the Lofdal mine site are 20°21′S 14°45′E. (WGS84)

Northern Namibia is a semi-arid environment. The property is characterised by gently rolling topography and is lightly forested. There is good road access to the property, and the town of Khorixas is connected to the national electricity infrastructure.

There is limited existing project infrastructure consisting of an exploration camp with PV power supply, a network of gravel roads and several water boreholes.

1.3. Geology and Mineralization

The Lofdal property is underlain by Paleoproterozoic metamorphic rocks of the Huab Metamorphic Complex, which outcrop as an inlier of the Congo Craton surrounded by stratified rocks of the Damaran Orogen. The metamorphic basement was intruded at ca 750 Ma by alkaline silicate rocks and carbonatites of the Lofdal Carbonatite Complex. The complex comprises an early silicate intrusive assemblage of dominantly nepheline syenite, and a later carbonatite intrusive assemblage ranging from calcitic through dolomitic and ankeritic carbonatites.

The Lofdal Carbonatite Complex comprises a central intrusive core characterized by several plugs of nepheline syenite and carbonatite with associated diatreme breccias, surrounded by a wide area of dyke intrusion and associated hydrothermal alteration. The phonolite and carbonatite dykes have exploited pre-existing structures in the basement that were re-activated during Neoproterozoic tectonism.

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Rare earth element mineralization in the Lofdal Carbonatite Complex is closely associated with the carbonatite dykes and related hydrothermal alteration. These occur within an area of more than 200 km². The lithogeochemical database demonstrates that many of the dykes are geochemically anomalous in REE (which includes yttrium as a heavy rare earth) with a significant number being of economic interest. Of particular significance is the frequent enrichment of heavy rare earths in the dykes and in structurally controlled hydrothermal alteration zones, which trend predominantly in NE - SW and NNE - SSW directions.

The rare earth elements (REE) are subdivided into heavy rare earth elements (HREE) and light rare earth elements (LREE). Lanthanum, cerium, praseodymium, neodymium, promethium, and samarium are the LREE. Yttrium, europium, gadolinium, terbium, dysprosium, holmium, erbium, thulium, ytterbium, and lutetium are the HREE. Although yttrium is lighter than the light rare earth elements, it is included in the heavy rare earth group because of its chemical and physical associations with heavy rare earths in natural deposits.

The rare earth mineralization in the Lofdal Carbonatite Complex is variable and includes both LREE and HREE enriched varieties that appear to have been introduced in separate mineralizing events. Petrographic evidence suggests that the heavy rare earth-rich mineralization resulted from a dominantly hydrothermal event that occurred relatively late in the history of carbonatite emplacement. Mineralogically, the heavy rare earth-enriched mineralization is dominated by xenotime (Y-REE phosphate), which is commonly associated with zircon, rutile, apatite and/or thorite. The mineralized hydrothermal alteration systems are continuous both along strike and at depth and produce clear geological, geochemical and radiometric signatures that are easily recognized, particularly in drillhole core.

1.4. Exploration Status

The Lofdal project was drilled by 411 boreholes for a total of 58,039 m.

The first Mineral Resource estimate in accordance with NI 43-101 was reported in 2012 for Area 4 based on geochemical analyses and density measurements of core samples obtained from 93 diamond drillholes completed by NMI in 2011 and 2012. An additional 17 diamond drillholes were drilled in 2012 and 2013 and a further 56 were drilled in 2020. In 2023, an in-fill RC drilling campaign was completed for Area 4 which added an additional 44 holes to the project resulting in an increase in confidence in the estimates and extension of the Mineral Resource at depth. For Area 2B, 17 diamond drillholes were drilled in 2010 and 2011 with an additional 29 drillholes completed in 2020. An additional 12 RC holes were drilled in 2023 which extended the Mineral Resource along strike in a northeasterly direction.

Drilling was orientated in a north to north-northwest direction with inclinations from 60° to 68° for Area 2B and from 55° to 71° for Area 4. Drillhole spacing for Area 4 is variable, with holes drilled in 2011 to 2012 positioned as close as 25 m apart. The 2020 campaign extended the Mineral Resource westwards with a 50 m spaced grid near surface, widening to 100 m down-dip. The 2023 RC campaign reduced this spacing, resulting in a staggered, 70 m spaced grid. At Area 2B, drillhole spacing is predominantly 50 m, widening to 100 m in the northeast strike extension. The RC drillholes were collared 100 m apart in the northeast part of the deposit with a single line of holes spaced 50 m apart in the southeast, which targeted the shallow, mineralised zones. Drilling has demonstrated that the mineralization continues down to a vertical depth of at least 380 m for Area 4 and 230 m for Area 2B.

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Table 1 Summary of drilling conducted at the Lofdal Project

Project Area	Drill Program	Type Drilling	2008-2016		JOGMEC 2020-2025		TOTAL PROJECT
			Holes	Meters	Holes	Meters	Holes	Meters
2, 2A, 2C	Reconnaissance	Diamond	13	1,265			13	1,265
2B	Resource	Diamond	17	2,134	29	4,400	46	6,534
2B	Resource	RC			12	1,780	12	1,780
2B	Geotech	Diamond			6	563	6	563
4	Resource	Diamond	110	12,635	56	10,162	166	22,797
4	Resource	RC			44	9,043	44	9,043
4	Metallurgy	Diamond	8	1,022			8	1,022
4	Geotech	Diamond			13	2,032	13	2,032
5	Reconnaissance	Diamond	57	5,595			57	5,595
6	Reconnaissance	Diamond	24	4,495			24	4,495
7	Reconnaissance	Diamond	1	240			1	240
8	Reconnaissance	Diamond	7	1,021			7	1,021
Northern Splay	Reconnaissance	Diamond			10	1,276	10	1,276
Dolomite Hill	Reconnaissance	Diamond			4	377	4	377
	TOTAL		237	28,407	174	29,633	411	58,039

Starter pit at Area 4 and bulk sample extraction

The Company developed an open pit in the central part of the Area 4 deposit. A first box cut of 60 m x 20 m to 15 m depth was excavated in 2022 and 30,000 t of material removed. A blended ore sample of 550 t was produced with a grade of 0.18% TREO and samples were sent to TOMRA (Hamburg, Germany) and Rados (Johannesburg, South Africa) for sorting tests. Additional samples were sent to Geolabs (South Africa) for geotechnical tests and to SGS Canada Inc. in Lakefield, Ontario ("SGS Lakefield"), for pilot-scale flotation and hydrometallurgical test work.

Figure 1 Lofdal Area 4 pit as of March 2025

A significant extension and deepening of the open pit took place in February 2025 (Figure 1) and a total of 15,000 t of material was excavated to a depth of 17 m. A total of 500 t bulk samples from five different ore zones were selected and crushed and screened. Three different bulk samples were prepared representing the hanging wall zone, main ore zone and footwall zone for bulk XRT and XRF sorting tests and subsequent flotation tests.

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1.5. Mineral Processing and Metallurgical Testing

Ore Sorting

Mineralization at Lofdal is amenable to XRT sorting by detection of higher density minerals which are co-genetic with xenotime. Results indicate that XRT sorting technology can provide significant upgrades to the run of mine ("ROM") by rejecting waste in form of albitite, gneisses, muscovite and chlorite schists.

Ore sorting tests were part of the company's value engineering during the PFS process "Lofdal 2B-4". The tested flowsheet for the PFS aimed at upgrading a low-grade stream by XRT sorting prior to flotation, with high-grade ore supplied directly to flotation.

Initial tests with TOMRA's new AI based and deep learning application OBTAIN, yielded improved XRT sorting results as compared to previous test programs. This formed the basis for bulk sample test programs carried out by Gecko Namibia with a full size TOMRA XRT sorter at the Ondoto LREE Mine in northern Namibia. The pilot-scale XRT test program was conducted on about 300 tons of ROM ore in July-August 2025. Sorting tests were conducted separately on bulk samples from the hanging wall, the main ore zone and the footwall zone as these three zones are characterized by different host lithologies (gneisses, pegmatites, amphibolites) and mineralization pattern.

A total of two hundred different test runs were conducted on a TOMRA COM Tertiary XRT at Gecko Namibia's facilities. The test work was conducted as a combination of two different image processing methods, Dual Energy and Inclusion Detection. A special Multi Density Class Model was applied to distinguish between six different sensitivities. For the inclusion detection TOMRA's deep learning-based classification CONTAIN™ was tested to detect visual patterns and textures to recover finely disseminated mineralization within the low contrast Lofdal material. The test results were steadily improved through 27 test settings by systematically adapting the multivariate test principles, parameters and algorithms based on the results as the program advanced. While the nature of mineralization with fine veins of xenotime is not the ideal type of material for XRT sorting, the test results exceeded the targeted upgrade and recoveries. The overall test results on low-grade (0.10-0.17% TREO) footwall and hanging wall ore yielded REE upgrade factors of 2.3 to 2.7 and REE recoveries of 60% to 70%.

Flotation test program

Flotation is the key step in beneficiation of Lofdal's xenotime ore. Flotation test work was carried out at SGS Lakefield and other international laboratories with over 170 individual flotation tests using several types of collectors, depressants and considered thrifting of physical flotation conditions. The Company expanded on SGS Lakefield's extensive experience in mineral processing of rare earth deposits. Earlier test programs compared upgrades and recoveries of XRF and XRT products through direct flotation followed by magnetic separation, and through magnetic separation followed by flotation. The test program was further amended to include flotation tests directly on the fresh, low-grade samples representing future run-of-mine grades.

The impact of high intensity conditioning ahead of flotation yielded improved flotation performance. Best flotation results regarding upgrade, recoveries and operating costs were achieved using moderate dosages of the collector Florrea 3900 and Calgon as depressant. Considering the better performing flotation tests, cleaner flotation concentrates yielded overall mass pulls of 2.7-3.9% with a product grade of 4-6% TREO and a recovery of up to 70% TREO. More importantly, the high value Heavy Rare Earth Elements, mainly hosted in xenotime, showed significantly better recoveries (58-75% HREO) than the Light Rare Earth Elements (49-58% LREO). Four bulk flotation tests demonstrated repeatable flotation performances on the low-grade feed material. The cleaner flotation produced a concentrate ranging from 4.7 - 6% TREO.

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The objective of the 2023-2024 test program was to scale up tests, including locked-cycle testing for a high level of confidence in metallurgy, and confirmation of engineering design criteria for PFS capital and operating cost estimation. The locked cycle tests were completed and confirmed a steady circuit. No significant detrimental effect was observed due to the recirculation.

A 5 tons run of mine ore sample, at head grade of 0.18% TREO, was shipped to SGS Lakefield for pilot plant testing in a continuous milling and flotation regime for recovery of a rare earth concentrate. The main objectives were to validate the flowsheet that had been developed at bench scale in a continuous pilot plant and to generate a sufficient mass of flotation concentrate for downstream hydrometallurgical test work. A continuous flotation pilot plant was run on the ROM Bulk-1 sample at an average throughput of 44 kg/h, for a total of about 105 hours of operation. The results of the flotation pilot plant closely matched the benchmark results and demonstrated the viability of the flowsheet in a scaled up and continuous operation. The total rare earth recovery in the second cleaner concentrate was 55.5% at a grade of 2.65% TREO (including yttrium) and an average mass pull of 3.8%. The average recoveries of terbium and dysprosium were 55.2% and 56.2%, respectively.

Flotation variability testing was conducted on sixteen individual samples that varied with xenotime mineralisation and contents of the gangue minerals were found to varied across the sixteen variability samples. Individual flotation testing of the variability samples was carried out, followed by testing on two composite samples. Two composite samples were made up from each of the respective areas with Area 2B, made up with slightly higher HREO portion as well as higher MgO and CaO and Area 4 samples which closer resembled the ROM composite head grade. TREO Recoveries ranged between 43% to 73% at grades between 1.5% to 4% TREO (estimated based on Yttrium). Varying reagent dosages had an effect on flotation performance with that of sodium silicate, Calgon and Florrea collectors critical. The flotation tests performed better at the finer grind size, minus 38 µm versus minus 53 µm.

Further flotation work continued in 2025 based on the recommendations from the variability work to investigate alternative depressants and collectors along with reagent make-up procedures plus further investigation of flotation performance at finer grinds on remaining ROM bulk sample.

Following the XRT ore sorting program on bulk blasted samples, upgraded ore sorted samples were tested for flotation performance. These comprised three ore sorted samples from footwall, hanging wall and main zone. TREO grades of the three ore sorted samples were 0.28%, 0.40% and 0.23% respectively. The response of the three sorted samples to flotation was reasonably good with minor modification to the reagent dosages. Furthermore, two bulk flotation tests using 10 kg charges were performed on the combined sample (Comb SP-2025) made from the three sorted samples. Test CP-202 results were better than the small batch flotation tests performed on the individual sorted samples. Recoveries ranged from 57 to 77% TREO, at lower concentrate grades between 1.5 to 1.9% TREO at higher mass pulls up to 5.9% as compared to benchmark test of 2.2%.

Hydrometallurgical test work and results

The PFS test program has shown that a simplified acid bake and liquor treatment flow sheet consisting of a high temperature acid bake, two stage (primary and secondary) impurity removal, followed by UIX and two stages (primary and secondary) of HREE carbonate precipitation is able to produce a high grade HREE carbonate. The flow sheet developed in this program has eliminated several unit operations from the original flowsheet in the PEA (Figure 2). The removal of crude REE precipitation, re-leach and thorium solvent extraction forms a significant simplification and overall reduced reagent demand.

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The following conclusions summarize key findings of the test program:

Under optimum operating periods, continuous high (600°C) temperature sulfation in a pilot rotary kiln yielded high HREE dissolution (94% Tb/Dy).
Batch tests were used to show that two stages of impurity removal using magnesium carbonate was able to remove practically all thorium, scandium, iron, aluminium and some of the uranium at minimum losses of HREE (~2%).
Uranium was removed by ion exchange using a conventional strong base anion resin (Puromet MTA4601PF). Uranium levels were reduced to below detection limit (0.02 mg/L U) with negligible co-extraction of HREE.
The U IX barren liquor was used in a mini pilot plant where a HREE carbonate was produced. The circuit consisted of two stages (primary and secondary) of precipitation using sodium carbonate.

Figure 2 Simplified hydrometallurgical flowsheet developed during PFS

Destruction of lixiviant and subsequent neutralisation with Magnesium Carbonate, is costly in hydrometallurgical flowsheets. Opportunities for acid optimisation and magnesium carbonate reduction should be further investigated.

1.6. Mineral Resource Estimate

Lofdal was visited by Jeremy Witley, who is a Principal Mineral Resource Consultant with MSA and the Qualified Person for this Mineral Resource Estimate, from 28 to 30 October 2020 on 10 November 2022 and from 21 to 22 November 2023. The occurrences and setting of the REE mineralization were observed in the field as well as the drilling in progress at the time. The mineralization was examined in a selection of drillhole cores from the 2020, 2022 and previous drilling programs. The QP was satisfied that the procedures and protocols used in drilling are consistent with CIM Exploration Best Practices Guidelines.

The assay results received from the primary laboratory (Actlabs in Ancaster, Ontario, Canada) were subjected to a quality assurance and quality control program and the assays have been confirmed by check assays completed by a second laboratory (ALS Minerals,

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North Vancouver, Canada). The drilling, logging, sampling and assay data are contained in a well organised drillhole database that the QP considers to be suitable for the purposes of mineral resource estimation.

The Mineral Resource Estimate was based on sample assays and density measurements obtained from the cores of diamond drillholes completed in three phases of drilling; 2011 to 2012; 2020 and the recent 2023 RC drilling campaign.

For the purposes of creating a framework for mineral resource estimation, fifteen mineralised zones were modelled for Area 4 and nine for Area 2B using a statistically defined cut-off of 10 ppm Dy₂O₃ for Area 4 and 12 ppm Dy₂O₃ for Area 2B. The resultant vein-like bodies within each deposit tend to be orientated parallel to one another, some of which coalesce in places at depth and along strike.

Ordinary kriging was used to estimate the individual rare earth element grades into a three-dimensional block model. Density was estimated into the same block model using inverse distance weighting. The Mineral Resource for Area 4 extends up to 1,600 m along strike near surface and attains a maximum depth of approximately 400 m. For Area 2B, the Mineral Resource extends for 850 m near surface and attains a maximum depth of approximately 230 m.

The Mineral Resource Estimate was completed by Mr. R. Goncalves (BSc Hons, MSc (Eng.)) under the supervision of Mr. J.C. Witley (BSc Hons, MSc (Eng.)).

The Mineral Resource was estimated using The Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Best Practice Guidelines (2019) and is reported in accordance with the 2014 CIM Definition Standards, which have been incorporated by reference into National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101). The Mineral Resource is classified into the Measured, Indicated and Inferred categories for Area 4 (Table 2) and into the Indicated and Inferred categories for Area 2B (Table 3)

The Mineral Resource is reported from a Whittle optimised pit shell at a base case total rare earth oxide (TREO) grade of 0.10%, which the QP considers will satisfy reasonable prospects for eventual economic extraction.

Table 2 Area 4, Measured, Indicated and Inferred Mineral Resource Estimates above 0.1% TREO cut-off grade – 05 April 2024

Category	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy₂O₃ ppm	TREO (kt)
Measured	6.6	0.21	0.14	0.07	130	13.7
Indicated	49.2	0.15	0.07	0.08	69	75.7
Measured & Indicated	55.8	0.16	0.08	0.08	76	89.4
Inferred	10.5	0.14	0.06	0.08	58	15.0

Notes:
1. All tabulated data have been rounded and as a result minor computational errors may occur.
2. Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
3. Quantities reported are the total quantities for the project regardless of ownership.
4. TREO = Total Rare Earth Oxides and includes Y₂O₃
5. HREO = Heavy Rare Earth Oxides and includes Y₂O₃
6. **LREO = Light Rare Earth Oxides
7. Mt = Million tonnes, kt = Thousand tonnes.

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Table 3 Area 2B, Indicated and Inferred Mineral Resource Estimates above 0.1% TREO cut-off grade – 05 April 2024

Category	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy₂O₃ ppm	TREO (kt)
Indicated	2.7	0.16	0.09	0.07	97	4.4
Inferred	4.4	0.15	0.07	0.08	75	6.6

1.7. Mineral Reserve Estimate

The Lofdal Rare Earths Project consists of two separate mineral deposits.

Area 4: (A4) Currently identified as the main mineralised area
Area 2B: (A2B) The second mineral deposit located approximately 2.5 km WNW of Area 4

Table 4 Lofdal Mineral Reserves as of 01 December 2025

Reserve Category	Mineral Deposit	Tonnes	Rare Earths Grade			Contained Rare Earths Oxide
			LREO	HREO	TREO	LREO	HREO	TREO
		(Mt)	(%)	(%)	(%)	(t)	(t)	(t)
Proven	Area 2B	-	-	-	-	-	-	-
	Area 4	6,19	0,068	0,144	0,211	4 194,0	8 893,2	13 087,1
Total Proven		6,19	0,068	0,144	0,211	4 194,0	8 893,2	13 087,1
Probable	Area 2B	1,90	0,075	0,094	0,169	1 430,3	1 792,8	3 223,1
	Area 4	23,91	0,076	0,091	0,167	18 269,3	21 761,6	40 030,7
Total Probable		25,81	0,076	0,091	0,168	19 699,7	23 554,4	43 253,8
Total Reserves		32,01	0,075	0,101	0,176	23 893,7	32 447,5	56 340,9

Notes on the Mineral Reserve:

Mineral reserves are reported at a cut-off grade of 0.100% TREO, based on a basket rare earths oxide price of USD 86.84/kg.
Reserves are based on open-pit mine designs with an average strip ratio of 6.8:1.
Metallurgical recovery at the hydrometallurgical plant is assumed at 62.2% for HREO and 52.55% for LREO.
Mass recovery for the >10mm primary crushed Low-Grade is assumed at 80%
Parameters at XRT Sorter is assumed at mass-pull of 25% and metal recovery of 65%
The reserve estimate was prepared by a Qualified Person (QP) in accordance with NI 43-101 and CIM Definition Standards (2014).
Mineral reserves are inclusive of mineral resources.

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1.8. Mining Methods

All the mine planning work for the PFS was done with Hexagon's MinePlan® (previously called MineSight) software. The proposed mining method is conventional open pit mining. Mineralised rock and waste would be drilled, blasted, loaded by hydraulic shovels and hydraulic excavators into off-highway dump trucks, and hauled to the processing plant. The proposed mining sequence the development of a slot-ramp along strike. This will enable selective waste mining on both sides of mineralised zones.

The basis for the pit design work was the mineral resource block model that was developed by MSA as part of a NI 43--101 mineral resource estimate (refer to Section 14).

There are two sub-deposits currently under consideration for mining. These are to be mined as open pits, with the normal sequence of drilling, blasting and hauling. Due to the nature of the deposit, the resultant pits are relatively narrow along strike and deep. Currently no backfilling is contemplated.

Variably sub-blocked Datamine geological block models were provided by the client as inputs. The Datamine block models were imported in Hexagon MinePlan software and verified against the unconstrained block models statistics reported by MSA.

The sub-blocked models were regularised for a selective mining unit (SMU) size of 5,0m x 5,0m x 2,5m for 62,5m³ (≈175 tonnes) for mine planning purposes.

The pit optimisation for the two pits was accomplished using Hexagon's MinePlan Project Evaluator (MPPE) task suite. The summary input data for the pit optimisation is shown in Table 5.

The target total ROM feed for processing is 3 010 000 tonnes/annum. On an annual basis, 1,91 Mtpa Low-Grade (0,10% ≤ TREO% < 0,16%) ROM is fed to the XRT ore sorter after primary crushing and screening. Additionally, 1,10 Mtpa High-Grade (TREO% ≥ 0,16) ROM is fed to the crushing and milling circuit and then to the flotation and hydrometallurgical plant. The upgraded ore concentrate from the XRT sorter amounts to 0,38 Mtpa and joins the High-Grade ore stream before secondary crushing and milling.

Table 5 Summary Pit Optimisation Parameters

Parameter	Units	a2b	a4
Commodity Price
Base TREO Basket Price	USD/kg	86,8352	86,8352
Base HREO Basket Price	USD/kg	131,8303	131,8303
Base LREO Basket Price	USD/kg	45,2896	45,2896
Selling Costs
Government Royalty	%	3,00%
Private Royalty	%	2,00%
Government Export Levy	%	0,00%
Selling Cost (Transport)	USD/t product	36,3100
Mining Operating Costs
Steady-state Ave Mining Cost	USD/t	2,9489	2,8499
Ref Ore Mining Cost	USD/t	2,8000	2,8000
Ref Waste Mining Cost	USD/t	2,8000	2,8000
Ore Depth Factor L&H	USD/t/5m-bench	0,0100	0,0100
Waste Depth Factor L&H	USD/t/5m-bench	0,0100	0,0100
Metallurgical Parameters
Process Recovery (TREO)	%	56,65%	56,38%
Milling Limit
High-Grade + Low-Grade	t/a	3 010 000
Process Cost

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Parameter	Units	a2b	a4
Average Process Cost	USD/t MILL Feed	29,2260	29,2260
High-Grade Process Cost	USD/t MILL Feed	28,8200	28,8200
Low-Grade Process Cost	USD/t MILL Feed	39,9650	39,9650
Incremental Ore Cost	USD/t ROM ore	0,2000	0,0000
Fixed Cost
Management - G&A	USD/a	3 045 600
	USD/t ROM ore	1,0118

The optimum pit shells were selected based on the maximum NPV bases. The pit designs followed the selected optimum pit shells templates to design the practical pit boundaries and in-pit ramps. The interim pushbacks (phases) for Pit A4 only were designed after the final pit to always ensure continuous bench-access between the phases as the pit are being mined.

The design process also includes the waste rock dumps for each mining area, topsoil stockpile locations, marginal stockpiles and ROM stockpiles.

The pit designs were split into mining cuts of approximately 130 000 tonnes and transferred to the MinePlan Schedule Optimiser for production scheduling. The main target for scheduling was to meet the plant feed on 3,01Mtpa with 1,91 Mtpa Low-Grade and 1,10 Mtpa High-Grade. The bench drop-down rate in every phase is also restricted to maximum 6-benches per year.

The resultant Life of Mine for the two pits is approximately 13 years, inclusive of pre-stripping and ramp-up activities. All cost and revenue modelling figures have been converted to US dollars. The results from the mining schedule includes all mining volumes to various destinations, the plant feed summary volumes and grades, stockpile rehandling, and load and haul mining equipment requirements. The output from the mining schedule was then transferred to the financial model for OPEX and CAPEX modelling.

1.9. Recovery Methods

SGS Bateman provided the detailed plant layout. The plant will consist of three main sections: the crushers and sorters, the concentrator and the refinery. The crushers and sorters will comprise ROM feed reception, low grade ore primary and secondary crushing, ore sorting and tertiary crushing while high-grade ore undergoes only primary, secondary and tertiary crushing.

The concentrator will comprise crushed ore stockpile, milling, and flotation. The flotation circuit will include roughers, cleaners, concentrate thickening and filtration, tails thickening and transfer of the tailings underflow to the tailings storage facility. The water recovered from the tailing and concentrate overflow thickeners will be pumped into the process water storage facility. The concentrate cake will be transferred to the refinery section.

The refinery section of the plant will be dedicated to the extraction, purification, and precipitation of REEs through sulphation roast, water leach, impurities removal, and REE precipitation. The REE will be precipitated as a Mixed Rare Earths Carbonate and packaged for sale.

1.10. Project Infrastructure

Project infrastructure will exist in the vicinity of the two open pits: A4 and A2B general site and the Mill Process Plant site.

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Water supply: The planned Lofdal plant requires about 1.5 million cubic meters of water annually. The Company drilled 16 groundwater boreholes in an aquifer system around Fransfontein, located about 35 kilometers to the northeast of Lofdal. SLR Namibia conducted pump testing on 10 boreholes. Six selected high yielding boreholes have a combined 48-hour Constant Discharge Test (CDT) yield of 237 m³/h, and a recommended abstraction of 180 m³/h which translates to an annual water supply of between 1.7 Mm³ and 1.3 Mm³. CREO Engineering Solutions (CREO) estimated the required CAPEX for the abstraction, delivery and site storage infrastructure at USD10 million.

Electricity supply: The Lofdal plant is expected to require approximately 94,361 MWh of electricity annually. CREO modelled the preferred bulk power supply mix consisting of grid connected power, supplied through the national power utility NamPower, supplemented by a third of the energy requirements through renewables (solar photovoltaic). The grid connection will require the construction of a new 200 km long 132 kV transmission line together with a 132/11 kV (20 MVA) main incoming substation at the project. The estimated CAPEX for the bulk power infrastructure is USD29 million.

Figure 3 Layout of planned mining and processing infrastructure at Lofdal

Road infrastructure: The mine will be connected to the main road network via road D2625 and a newly constructed 10 km gravel road connecting the plant area.

Tailings Storage Facility: Effluent streams from the above section will be pumped into the neutralization tank from where it will be transferred with the tailings to the tailings storage facility ("TSF"). The TSF footprint accommodates potential future mine life extension and long-term storage requirements.

A total of approximately 16 million tons of tailings are expected to be produced over the Project Life-of-Mine. Tailings thickened at 46% solids content by mass will be pumped and conveyed to the TSF located east of the Process Facility and the Area 4 Main Open Pit (see Figure 3). The TSF comprises a cross valley compacted earth fill starter embankment with liner system over the embankment upstream face and basin, covered with an underdrainage system to increase tailings dewatering and water recycle.

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The TSF construction strategy includes an initial downstream raise using selected waste rock from the open pit placed during the first years of operation, followed by an upstream raising strategy to final elevation. The TSF final height is approximately 28 m, and the embankment was sized to accommodate potential future expansion inside the valley. Tailings will be discharged through spigots along the face of the embankment and side hills for pool control. Decant water will be pumped back to the processing facility from the return water dam. Tailings classify as silt with trace clay and are non-acid generating with no neutralisation potential.

A liner system is included in the PFS design to reduce seepage and water losses through the weathered foundation. Further geotechnical and hydrogeological investigations along with radionuclide testing will be required to design the most appropriate seepage mitigation measure in the next design stage and potentially reduce the requirement for a liner system while maximizing seepage recovery. The design includes a provision for monitoring instruments such as piezometers and level control system as well as dust mitigation through progressive capping of the TSF.

1.11. Rare Earth Pricing

A price deck has been developed for the Lofdal Project based on an independent forecast provided by CRU International Limited ("CRU"), Argus Europe assessments and publicly available third-party intelligence.

The rare earth market is characterized by a bifurcation between China-domestic and ex-China prices for Nd, Pr, Dy, Tb and Y. Export controls and security-of-supply concerns are creating a distinct ex-China price regime for critical raw material deriving from sources outside China. Lofdal's future products are expected to participate in this ex-China price environment.

Divergence price trends

Export restrictions and supply strategy has created a "China price" and a higher "rest-of-world price" for the same material.

Dy/Tb
Chinese spot prices for Dy and Tb oxides (e.g., Dy oxide ~USD240/kg, Tb oxide~USD1,000/kg) represent domestic or FOB China values.
According to Benchmark Minerals, markets for heavy rare earths have been facing significant pressure since April 2025 with actual spot prices for ex-China supply reaching USD900/kg for Dy oxide and USD3,625/kg for Tb oxide for imports into the European Union.
Long-term assumptions used in recent PFS/DFS studies effectively embed an ex-China premium driven by supply-chain security (e.g. Carina uses USD829/kg Dy oxide and USD3,056/kg Tb oxide).
Nd/Pr
NdPr remain global commodities with relatively tight arbitrage between China and ex-China, but the same forces (tariffs, export controls, strategic stockpiling and ESG filters) are pushing contract prices for ex-China Nd/Pr feedstock above Chinese spot prices for long-term secure supply.
This has been reinforced with the announcement of the US Department of War investing in MP Materials and establishing a floor price of USD110/kg for NdPr oxides, almost double the China market pricing at the time.

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Yttrium
The Lofdal Mixed Rare Earths Oxide product contains about 40 to 50% Yttrium oxide.
European spot prices for yttrium oxide have risen as much as 4,400% since January 2025 up to USD270/kg.
For the medium to long term (2026–2041), the Base Case assumption is that export controls ease partially and new ex-China supply gradually ramps, resulting in an indicative yttrium oxide price range of USD30–80/kg (real 2025 dollars).
A tight-supply case (Divergent Case), reflects prolonged export restrictions or slow non-Chinese supply growth, and places yttrium oxide in the USD80–150/kg band, consistent with recent ex-China price behavior.

These pricing bands are planning assumptions and do not constitute market forecasts.

The rare earth oxide pricing used in the PFS (average Life of Mine) for the three main value drivers are:

Base Case: Dy2O3 USD663/kg, Tb2O3 USD2,880/kg and Y2O3 USD60/kg
Divergent Case: Dy2O3 USD855/kg, Tb2O3 USD3,712/kg and Y2O3 USD130/kg.

$Y_{2}O_{3}$ represents approximately 40–50% of Lofdal’s recovered rare earth oxide basket.

1.12. Socio-Economic and Environmental Impact

The anticipated impacts associated with the Project were assessed according to SLR’s standardised impact assessment methodology. The impact assessment methodology enables the assessment of biophysical, cultural, and socio-economic impacts including cumulative impacts and impact significance through the consideration of intensity, extent, duration, and the probability of the impact occurring.

The assessment indicates that the Project’s impacts range from highly positive to highly negative. Notably, the assessment assumes that the group of residents currently residing near the Project (specifically at Oas Post 3 and the Lofdal Homestead, within the vicinity of the Area 2B WRD and TSF) will be relocated prior to the commencement of mining construction activities.

A comparative analysis of the impacts assessed in the 2016 EIA indicates that the revised Lofdal Mining Project will not result in increased adverse impacts. The changes in layout and LOM have led to an expanded operation with an extended LOM, contributing to impacts with a higher positive significance rating. While certain potential impacts identified during the Impact Assessment Phase were classified as highly significant, appropriate mitigation measures, as outlined in the EMP, can effectively minimize these adverse effects. Implementation of the EMP will be subject to ongoing monitoring and auditing to evaluate the efficacy of the prescribed mitigation measures.

All negative environmental and social impacts identified will be managed and mitigated to acceptable levels, whilst the positive impact will be enhanced to realise the potential positive impacts through the implementation of the commitments stipulated in the EMP. NRE will be responsible for ensuring that all environmental and social obligations pertinent to the proposed Project are met. The implementation of the EMP and meeting of the environmental objectives and targets are Lofdal Mining has been prepared and attached in Appendix H. The EMP contains specific management measures recommended by the specialists that should be implemented.

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It is anticipated that it will be possible to successfully mitigate all the environmental impacts to acceptable levels and the implementation will be monitored and audited to determine the effectiveness of the measures implemented.

No fatal flaws/aspects have been identified that could render this the Project unfeasible and impractical. Therefore, it is SLR's opinion that, based on the findings of the EIA process, there is no reason why the proposed development may not continue subject to the implementation of recommended mitigation measures. The Project should be allowed to proceed, considering the positive social and economic benefits associated with the Project.

1.13. Legal and Statutory

For the Project to proceed successfully, a number of legislative requirements will need to be fulfilled according to the Namibian legislation and possible international legislation and guidelines. NMI will be responsible for ensuring that the welfare of the local population is not significantly impacted upon due to the mining activities. In addition, NMI must ensure that adequate rehabilitation and closure of the mine takes place following the conclusion of the proposed mine.

To ensure that the legislative requirements are met, as well as best practices are implemented, environmental degradation and pollution must be prevented and, where unavoidable, mitigated, and managed. The predominant impacts associated with the mining activities are due to groundwater quantity, potential groundwater and dust contamination and the potential side effects of thorium. Other social ills may result from the project due to the influx of job seekers causing an increase in the population of Khorixas.

1.14. Capital and Operating Cost Estimate

Mining will be conducted via contractor, and all contractor capital recovery is reflected in the mining operating costs. A portion of the mining capital is for contractor mobilization, with the majority of capital applied to pit pre-stripping.

Process capital includes the process plant and ore sorting facility.

Facilities capital includes all non-process site facilities, including water and power supply, non-process site buildings, security and warehousing.

CAPEX increases reflect inflation since the 2022 PEA, expanded hydrometallurgical scope (acid recovery), inclusion of mining pre-strip and revised power infrastructure requirements.

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Table 6 Capital Costs Summary of Lofdal PFS "Lofdal 2B-4"

Capital Costs Summary (USD)
Mining Capital*	$ 27,620,316
Process Capital	$ 181,571,767
Facilities Capital	$ 58,746,582
Tailings Capital	$ 21,590,251
Closure Costs	$ 1,039,987
Sub-Total	$ 290,568,903
Contingency	$ 57,360,067
Total Capital Costs	$ 347,928,970

OPEX increases from the 2022 PEA are driven by higher acid and reagent prices, diesel kiln operation and updated power tariffs.

Table 7 Operating Costs Summary of the Lofdal PFS "Lofdal 2B-4"

Operating Costs Summary (USD)
	Life of Mine	Per tonne mined	Per tonne processed	Per kg TREO
Mining Cost	$ 652,439,644	$2.63	$37.32	$24.78
Processing	$ 996,304,067		$56.98	$37.84
G&A	$ 29,783,297		$1.70	$1.13
Total Operating Costs	$ 1,678,527,008		$96.00	$63.75

Royalties and separation costs are based on total gross revenue.

Table 8 Royalties and Separation Costs of Lofdal PFS "Lofdal 2B-4"

Other Pricing Model Costs – Life of Mine
	Base Case	Divergent Case
Royalties and Separation Costs	$ 295,383,503	$ 364,742,532

1.15. Economic Analysis

The economic analysis assumes that the Project will be 100% equity financed and uses parameters relevant as of December 2025, under conditions likely to be applicable to project development and operation and analyzes the sensitivity of the Project to changes in the key Project parameters. All costs have been presented in United States Dollars (USD) and wherever applicable conversion from South African Rand (ZAR) has utilized an exchange ratio (ZAR/USD) of 18.23.

Mining and treatment data, capital cost estimates and operating cost estimates have been put into a Base Case and Divergent Case financial model to calculate the IRR and NPV based on calculated Project after tax cash flows. The scope of the financial model has been restricted to the Project level and as such, the effects of interest charges and financing have been excluded.

For the purposes of the PFS, the evaluation is based on 100% of the Project cash flows before distribution of profits to the equity owners. Both pre-tax and after-tax cash flows have taken 5% royalty payments into account.

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In the Base Case, the Project is anticipated to yield pre-tax IRR of 21.7% and NPV of USD389,158,821 (using a discount rate of 5% in all cases), and after-tax IRR of 19% and NPV of USD275,510,605. In the Divergent Case, the Project is anticipated to yield pre-tax IRR of 44.1% and NPV of USD1,245,607,396, and after-tax IRR of 34.8% and NPV of USD747,870,616.

Cumulative cash flows over the 13 years mine life are: Base Case USD709,589,893 pre-tax and USD513,086,468 after-tax, and Divergent Case USD2,027,411,439 pre-tax and USD1,242,332,502 after-tax.

The Project is expected to pay back initial capital within the first 4.2 years (Base Case) and alternatively in 2.75 years (Divergent Case).

Sensitivity Analysis

Economic Sensitivities & Cash Flow Summary (After-Tax, 5% Discount Rate)

The Prefeasibility Study confirms that the Project delivers strong early cash flow, rapid capital recovery, and economic resilience under conservative pricing assumptions, with upside leverage under divergent rare earth pricing scenarios. Lofdal exhibits high sensitivity to yttrium pricing due to its HREE dominant basket.

After-Tax Cash Flow Profile

Under the Base Case, the Project generates consistent positive after-tax cash flows throughout the 13 years mine life following construction, with cumulative after-tax cash flow turning positive early in operations and increasing steadily to closure. Annual after-tax cash flows strengthen rapidly following ramp-up and remain stable across mid-mine and late-mine production.

Under the Divergent Case, reflecting higher rare earth pricing, the Project exhibits a material uplift in early-year cash flows and accelerated cumulative cash flow growth, with cumulative after-tax cash flow exceeding USD1.25 billion. This case highlights the Project's strong exposure to structural tightness in dysprosium, terbium and yttrium markets.

The Project achieves rapid capital payback of approximately 4.2 years after-tax in the Base Case and approximately 2.75 years in the Divergent Case, supporting strong financing attractiveness.

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Figure 4 Illustrative cash flow profile for Base Case and Divergent Case; year-on-year variability occurs in practice

After-Tax Price, Cost & Exchange Rate Sensitivities

The after-tax sensitivity analysis demonstrates that:

Metal prices are the dominant value driver, with a ±20% change generating the largest impact on NPV in both, Base and Divergent cases.
Operating costs represent the second-most influential variable; however, the Project retains a positive after-tax NPV across all tested cost ranges.
Exchange rate movements provide additional economic leverage, with a weaker local currency significantly enhancing project value.
Capital costs show moderate sensitivity, confirming that the Project's value is not disproportionately dependent on Capex precision.

Importantly, under the Base Case, the Project maintains positive after-tax NPV across all tested price, cost and exchange-rate sensitivity ranges, demonstrating strong downside protection. Under the Divergent Case, after-tax NPV expands materially under higher pricing and remains highly robust under adverse cost scenarios.

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Figure 5 Sensitivities for Base Case and Divergent Case

Process Recovery Sensitivity

Metallurgical recoveries represent a high impact, controllable value lever for the Project:

Under Base Case pricing, after-tax NPV increases from approximately USD100 million at 85% of expected recoveries to approximately USD440 million at 115% of expected recoveries.
Under Divergent Case pricing, after-tax NPV increases from approximately USD500 million at 85% of expected recoveries to approximately USD1.0 billion at 115% of expected recoveries.
The linear and consistent response of NPV to recovery improvements demonstrates that ongoing metallurgical optimisation provides meaningful value upside, while the Project remains economically viable even at materially lower-than-design recoveries.

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Figure 6 Process Recovery Sensitivities for Base Case and Divergent Case

Overall Economic Interpretation

The combined cash flow and sensitivity analyses confirm that the Project:

Is financially robust under conservative assumptions;
Exhibits exceptional leverage to critical heavy rare earth pricing, particularly dysprosium, terbium and yttrium;
Benefits from strong operating margin resilience to cost pressures;
Generates early and sustained after-tax cash flow, supporting attractive project financeability; and
Provides significant embedded strategic optionality in a tightening global heavy rare earth supply environment.

1.16. Conclusions

This PFS demonstrate that the Lofdal Heavy Rare Earth Project has the potential to be technically and economically viable as a producer of rare earth elements carbonate.

1.17. Recommendations

The PFS established a two-stream model as the ideal flow sheet for Lofdal, whereby higher-grade material enters flotation directly and lower-grade material will be upgraded by XRT sorting prior to flotation. Both material streams need constant supply which is challenging by operating one or two open pits only. To increase the efficiency of the upfront processing plant, the development of additional satellite pits as swing producers is advisable. These pits can supply additional supplementary high-grade or low-grade material as required during steady operation. Therefore, resource drilling of Area 5 is recommended.

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Current mine life is estimated at 13 years. The PFS made provisions for an extension of mine life by designing a significantly larger than required footprint of the tailings storage facility. Inferred Resources in the northern part of Area 2B are recommended for infill drilling to potentially add additional resources at higher resource categories. Further, reconnaissance drilling at Area 5 showed significant intercepts of HREE mineralization. With its location just to the north of the planned processing plant, it is ideally situated as potential swing producer.
Mine life can likely be significantly extended by underground mining of the extension of the ore zone at Area 4 at depth. Preliminary studies showed viable options of underground mining at Area 4. These studies should be elevated to DFS level.

The detailed recommendations list is provided in Section 27.

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2. Introduction

This Report was prepared and compiled by the QPs under employment with SGS and Consultants at the request of Namibia Critical Metals Inc. ("Namibia Critical Metals" or the "Company" or "NMI"). The purpose of this Report is to provide a Prefeasibility Study ("PFS") for the Lofdal Heavy Rare Earth Project "2B-4" ("Lofdal" or the "Project") in Namibia.

The Lofdal deposit has the potential for significant production of dysprosium ("Dy"), terbium ("Tb") and yttrium ("Y") which are the main economic drivers for the Lofdal project. The PFS presents a Base Case – representing a conservative but realistic post-normalization scenario, assuming partial easing of China's recent export bottlenecks and a Divergent Case - consistent with consensus divergence pricing logic and the geopolitical reality of prolonged export controls, slow non-Chinese separation build-out and rising strategic demand from OEMs.

This Technical Report has been prepared to comply with disclosure and reporting requirements set forth in the Toronto Venture Exchange (TSX-V) Corporate Finance Manual, Canadian National Instrument 43-101, Companion Policy 43-101CP, Form 43-101F1, the 'Standards of Disclosure for Mineral Projects' of January 2011 (the Instrument) and the Mineral Resource and Reserve classifications as defined in the CIM Definition Standards 2014 document.

Namibia Critical Metals (NMI) is a Canadian company listed on the TSX Venture Exchange and OTCQB Market which is developing the Lofdal Heavy Rare Earth Project within the Republic of Namibia. The subject of this technical report is the Lofdal Heavy Rare Earth Project 2B-4 (Lofdal) which is held in a Joint Venture Agreement with Japan Oil, Gas and Metals National Corporation (JOGMEC). The Company's registered corporate office is Suite 802, Sun Tower, 1550 Bedford Highway, Halifax, Nova Scotia, NS B4A 1E6 Canada.

2.1. Purpose of Report

The purpose of this Report is to publish a Technical Report on the Lofdal Heavy Rare Earth 2B-4 Project summarizing:

the land tenures, exploration history, and drilling;
the mineral resource estimates at "Pit 2B" and "Pit 4";
a conceptual mine plan at a level to support a Prefeasibility study;
the supporting infrastructure including, power, buildings, tailings management facility;
processing plant, etc. to support the mine plan;
the environmental permitting requirements;
capital expenditure and operating expenditure estimates;
a financial model and an economic analysis and;
provide recommendations and additional work.

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2.2. Terms of Reference

NMI engaged the services of SGS and authors in January 2024, to write an independent NI 43-101 Technical Report on the Lofdal Heavy Rare Earth Project 2B-4 Property in Namibia. This Report was prepared in accordance with NI 43-101 and Form NI 43-101 F1 and Companion Policy 43 101CP.

2.3. Qualifications of Consultants

SGS and Consultants preparing this technical report are specialists in the fields of geology, exploration, mineral resource estimation, open pit mining, geotechnical, environmental, permitting, metallurgical testing, mineral processing, processing design, civil, mechanical, electrical, capital and operating cost estimation, and mineral economics.

None of the Consultants or any associates employed in the preparation of this report has any beneficial interest in NMI. The Consultants are not insiders, associates, or affiliates of Namibia Critical Metal Inc. The results of this Technical Report are not dependent upon any prior agreements concerning the conclusions to be reached, nor are there any undisclosed understandings concerning any future business dealings between NMI, SGS and the Consultants. The Consultants are being paid a fee for their work in accordance with normal professional consulting practice.

2.4. Report Responsibility and Qualified Persons

The following individuals, by virtue of their education, experience, and professional association, are considered Qualified Persons (QP) as defined in the NI 43-101 standard, for this report, and are members in good standing of appropriate professional institutions:

MSA Group under the supervision of Jeremy Witley, (BSc Hons, MSc (Eng.)): Sections of the Report dealing with Property Description and Location (Item 4), Accessibility, Climate, Local Resources, Infrastructure and Physiography (Item 5), History (Item 6), Geological Setting and Mineralisation (Item 7), Deposit Types (Item 8), Exploration (Item 9), Drilling (Item 10), Sample Preparation, Analyses and Security (Item 11), Data Verification (Item 12) and Mineral Resource Estimate (Item 14).
SGS North America Inc. under the supervision of Joseph M. Keane, (B.S, M.S., PE Metallurgy): Sections of the Report dealing with the Mineral Processing and Metallurgical Testing (Item 13).
Qubeka Mining Consultants under the supervision of Peter Christians (FAusIMM): Sections of the report dealing with Mineral Reserve Estimates (item 15) and Mining Methods (Item 16).
SGS Bateman (Pty) Ltd under the supervision of Joseph M. Keane, B.S, M.S., PE Metallurgy: Section of the Report dealing with the Recovery Methods (Item 17).
CREO Engineering Solutions (Pty) Ltd under the supervision of Etienne Roux (Pr. Eng): section of the Report dealing with Project Infrastructure (Item 18 except for Items 18.10).
Knight Piésold Consulting (Pty) Ltd under the supervision of Veronique Daigle (Eng. / Pr.Eng.): Section of the Report dealing with Tailings Storage Facility (TSF) (Item 18.10).
SGS Canada under the supervision of William van Breugel (B.Sc. Hons, P.Eng.): Sections of the Report dealing with Market Studies and Contracts (Item 19),

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Environmental Studies, Permitting and social or community impact (Item 20), Capital and Operating Costs (Item 21), and Economic Analysis (Item 22).

The preceding QPs have contributed to the writing of this Report and have provided QP certificates, included at the end of this Report. The information contained in the certificates outlines the sections in this Report for which each QP is responsible. Each QP has also contributed figures, tables, and portions of Sections 1 (Summary), 2, (Introduction), 3 (Reliance on other Experts), 25 (Interpretation and Conclusions), 26 (Recommendations), and 27 (References).

2.5. Site Visit

Personal inspections made by the Qualified Persons and their items of responsibility for this report are shown in Table 9.

Table 9: Details of Site Visits and Responsibilities of the Qualified Persons

Qualified Person	Personal Site Inspection Dates	Items Responsible for	Items Co-Responsible for
Jeremy Witley	October 28 to 30, 2020; November 10, 2022; November 21 to 22, 2023	4, 5, 6, 7, 8, 9, 10, 11, 12, 14 and 23	1,2,3,24,25,26 and 27
Peter Christians	11&12 December 2025	15 and 16	1,2,3,24,25,26 and 27

2.6. Currency, Units, Abbreviations and Definitions

All units of measurement used in this technical report are International System of Units (SI) or metric. Every effort has been made to clearly display the appropriate units being used throughout the Report. All currency is in US dollars (USD or $), unless otherwise noted. The locations of all maps are referenced to WGS 84, UTM Zone 33S, unless otherwise stated. Frequently used abbreviations and acronyms can be found in Table 10.

This Report includes technical information that required subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, the QPs consider them immaterial.

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Table 10 List of Abbreviations

Abbreviation	Description
%	Percent sign
°	Degree
°C	Degree Celsius
cm	Centimetre
g	Grams
g/t	Grams per metric tonne
Ga	Billion years
GPS	Global Positioning System
ha	Hectare
ICP-MS	Inductively coupled plasma mass spectrometry
ICP-OES	Inductively coupled plasma optical emission spectrometry
kg	Kilograms
km	Kilometres
m	Metres
m³	Cubic metres
Ma	Million years
mm	Millimetre
MRE	Mineral Resource Estimate
Mt	Million tonnes
N, S, E, W	North, South, East, West
ppm	Parts per million
QA	Quality Assurance
QC	Quality Control
QP	Qualified Person
SG	Specific Gravity
SGS	SGS Canada Inc. Geological Services
SGS Lakefield	SGS Minerals Services Lakefield Facility
tonnes or t	Metric tonnes
μm	Micrometers
USD	US Dollar
UTM	Universal Transverse Mercator

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2.7. Effective Date

The effective date of this technical report is December 3, 2025.

As of the effective date of this Report, the authors are not aware of any material fact or material change with respect to the subject matter of this Technical Report that is not presented herein, or which the omission to disclose could make this Report misleading.

2.8. Previous Technical Reports

A Preliminary Economic Assessment (PEA) on the Project was completed by SGS Canada Inc. on 14 November 2022, titled “Preliminary Economic Assessment on the Lofdal Rare Earths Project, Namibia”. Information considered by the QPs to be both current and relevant was sourced from this document.

A Mineral Resource Estimate (MRE) on the Project was completed by MSA on 24 May 2024, titled “NI 43-101 Technical Report – 05 April 2024 Mineral Resource Estimate” Information considered by the QPs to be both current and relevant was sourced from this document.

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3. Reliance on other Experts

MSA/SGS has not independently verified, nor is it qualified to verify, the legal status of the Lofdal property. The report has been prepared on the assumption that the tenements will prove lawfully accessible for evaluation.

The QPs have reviewed and analyzed data and reports provided by NMI, together with publicly available data, drawing its own conclusions augmented by direct field examination.

The QPs who prepared this report relied on information provided by experts who are not QPs. The QP believes that it is reasonable to rely on these experts, based on the assumption that the experts have the necessary education, professional designations, and relevant experience on matters relevant to the technical report.

SGS has relied upon Arnold Bittner – SLR on matters pertaining to the Summary of Environmental Impact Assessment Report and special Studies for the Lofdal Mining Project dated March 2021 as disclosed in Section 20.

William van Breugel, P. Eng. (SGS) has relied upon NMI who supplied pricing forecast for this PFS and derived from recent market analysis and other published NI 43-101 complaint resource reports on selling prices for REE, as summarized in Section 19.

William van Breugel, P. Eng. (SGS) has relied upon Knight Piesold, who completed an independent analysis on the Tailings Storage Facility for the data used in the Capital and Operating Expenses estimate as summarized in Sections 21.

William van Breugel, P. Eng. (SGS) has relied upon Alva Shortt and SGS Bateman, who completed an independent analysis on the Process Plant quantities and costs for the data used in the Capital and Operating Expenses estimate, and Economic Analysis as summarized in Sections 21 and 22.

The QPs have assumed, and relied on the fact, that all the information and existing technical documents listed in the References Section 27 of this report are accurate and complete in all material aspects. While the QPs reviewed all the available information presented, we cannot guarantee its accuracy and completeness. The QPs reserve the right, but will not be obligated, to revise the report and conclusions, if additional information becomes known subsequent to the date of this report.

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4. Property Description and Location

4.1. Property Location

The Lofdal property comprises of Mining License (ML) 200 and is located approximately 25 km northwest of the town of Khorixas in the Kunene Region of northwestern Namibia. Khorixas is approximately 325 km in a straight line and 450 km by paved road northwest of the capital Windhoek.

Source: MSA (2024)
Figure 7 Location of the Lofdal Property (red square NW of Khorixas)

4.2. Property Description

4.2.1. Mining License (ML) 200

The Lofdal property was originally held under Exclusive Prospecting License (EPL) 3400 granted in 2005. EPL 3400 was relinquished in November 2023 and replaced by Mining Licence 200 (ML 200). The ML 200 boundary within EPL 3400 is shown in Figure 8.

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Source: Base is September 2022 Google Earth - UTM WGS84, Zone 33S. Compiled by NRI.
Figure 8 Location of ML 200 and the boundary of EPL3400 at the time it was relinquished showing current boundaries, roads, and the location of the Hoppe Mineral Claims and Mining License Applications

The mining licence application was lodged by Namibia Rare Earths (Pty) Ltd on November 16, 2016. Notice was received on December 22, 2020 that the Minister of Mines and Energy was prepared to grant the application for a mining licence. ML200 was granted on May 11, 2021 for a period of 25 years (expiring on 10 May, 2046) in respect of "Base and Rare Metals of Minerals" subject to certain terms and conditions (Ministry of Mines and Energy, Republic of Namibia, 2020), which are as follows:

“Part 1 - General

The mining licence shall endure for a period of twenty five years (25) reckoned from the date of acceptance (hereinafter "the date of issue") of the terms and conditions referred to in this notice unless it is abandoned in terms of section 54 of the Minerals (Prospecting and Mining) Act, 1993 (hereinafter "the Act") or cancelled in terms of section 55 of the Act or an application made to the Minister in terms of Section 96 of the Act, it is renewed by the Minister for any further period or periods.
In consideration of the rights hereby granted, the holder of the mining licence shall pay to the Commissioner for the benefit of the State Revenue Fund, such licence fee as may from time to time be prescribed in terms of section 123 of the Act, it being recorded that the annual licence fee prescribed in relation to the licence at the time of its issue shall be N$5 000.00 payable annually on or before each anniversary date of the date of issue of the licence.

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Part 2 – Work Program and Obligations

The holder of the licence shall:
1.1 commence with, and thereafter continue without undue interruption or delay, mining operations within one month of the date of issue of the licence in substantial conformity with the proposed work program, schedule and budget which accompanied the original application for the licence, and which served as motivation of the granting thereof;
1.2 where any material deviation of such work program, schedule and budget is in the opinion of the holder of the licence, necessitated by the nature of the results of mining operations (but specifically excluding any circumstances of Vis Major provided for in terms of section 56 of the Act), apply in writing to the Minister for approval of the revision of such work program, schedule and budget in terms of section 99 of the Act; and
1.3 execute such additional work program and expend such additional expenditure within a specified period of time as may be imposed by the Minister from time to time.
1.4 The Minister may, in the interest of the reasonable development of the mining operations, impose from time to time such additional terms and conditions as may deem fit.
2.1A all funds raised anywhere in respect of this licence shall be committed to this licence and shall be banked at a Financial Institution in Namibia.

Part 3 – Environment

The holder of the mining licence shall observe any requirements, limitations or prohibitions on his or her prospecting operations as may in the interest of the environmental protection, be imposed by the Minister.
The holder of the Exclusive Prospecting Licence shall adhere to the terms and conditions upon which the Environmental Clearance Certificate was issued by the Ministry of Forestry Environment and Tourism.

Part 5 – Additional Conditions

Within 30 days of the grant of a new Mining Licence, the applicant shall submit to the Minister a declaration signed by a duly authorised director of the applicant to the effect and including:
1.1 Proof that there is a minimum 20% representation of historically disadvantaged Namibians in the management structure (including the board) of the applicant; and
1.2 Proof that at least 5% (five percent) of the principal voting shares in the applicant or at least 5% (five percent) of the holding of the Mining, Licence, as the case may be, is held by historically disadvantaged Namibians. For the purposes of this condition, the term "held" includes a holding of such principal voting shares directly or indirectly through a company, close corporation, trust, traditional authority, or other similar association, and includes ownership by entities representing Government or in which Government holds a meaningful stake.
1.3 The applicant's strategy for addressing the Government's objective of poverty eradication, including benefitting the Namibian youth and women from disadvantaged groups and the poorest of the poor.
If the applicant has been misleading in relation to declarations made under condition 13, the Minister may cancel the licence under section 55(1)(a) of the Act and the remaining provisions of section 55 will apply.

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For the purposes of these conditions, the term “historically disadvantaged Namibians” shall mean Namibian citizens falling within the category of designated groups” as defined in the (Affirmative Action (Employment) Act, 1998).”

The Mining License area is shown in Figure 9 and Figure 10, and the coordinates of the boundary corners are given in Table 11.

Issued in favour of: Namibia Rare Earths (Pty) Ltd

Latitude and Longitude lines refer to the Bessel 1841 Spheroid

AREA: 21034.6274 Hectares

Source: Ministry of Mines and Energy, Republic of Namibia, 2020

Figure 9 Location of Mining License 200 (blue line). Corner numbers in red are same as given in Table 11.

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Table 11 Coordinates of Mining License 200
| Corner Point | Longitude (Degrees, Minutes, Seconds) | Latitude (Degrees, Minutes, Seconds) |
| --- | --- | --- |
| 1 (NE) | 14 49 9.99 E | -20 14 40.83 S |
| 2 (SE) | 14 49 9.92 E | -20 21 56.49 S |
| 3 (SW) | 14 39 10.06 E | -20 21 55.59 S |
| 4 (NW) | 14 39 10.68 E | -20 17 23.87 S |
| 5. (top) | 14 44 26.94 E | -20 14 38.08 S |

Note: The boundaries of the ML are established by reference to latitude and longitude coordinates in reference to the Bessel 1841 Spheroid, Central Meridian 17 degrees East

Source: Ministry of Mines and Energy, Republic of Namibia, 2020

Figure 10 Extent of Mining License 200 and the Former EPL 3400 in relation to the Mineral Resources of Area 4 and Area 2B

MSA has examined the documentation regarding Mining Licence 200 as supplied by NMI for review. Although MSA is not qualified to provide legal opinion, it has no reason to doubt the authenticity of the information provided.

4.2.2. General Provisions

Under the Minerals Act, 1992, and as declared in Government Gazette 45 of 2009, REE are subject to a royalty of three percent of the fair market value of minerals produced in Namibia. The property is also subject to a two percent Net Smelter Royalty (NSR) to Alberto Lobo-Guerrero Sanz, who introduced NMI to the project.

Neither the applications by NMI to acquire or renew the ML, nor the environmental contract that was agreed to by NMI and the Government of Namibia (Environmental Contract), identify any pre-existing environmental liabilities on the property and none are known to exist.

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The Environmental Clearance Certificate (ECC) for ML-200 was originally issued December 05, 2017 (for the Mining License application) and was subsequently renewed in 2021 and 2024. The latest renewal is dated April 24, 2025, and is valid until April 24 2028.

Notifications of trenching and drilling programs are required to be filed with the Mining Commissioner, Department of Mines and Energy. Notification of drilling for all holes in the 2020 drilling campaign were filed by forms dated February 26, March 13, July 8, August 12, September 24, and November 19, 2020. Additionally, August 30, 2021, October 22, 2022 and for the 2023 RC drilling campaign on January 20, and October 23, 2023. The authors are not aware of any other permits that are required to conduct the planned work.

4.2.3. Adjacent and Overlapping EPLs

The Lofdal Carbonatite Complex is entirely contained within ML 200. As far as is known, there are no similar intrusions or potential for similar mineralization outside the ML and there is no active exploration for similar targets on nearby EPLs.

The area of the former Lofdal copper mine is held under mining claims by a Mr. Hoppe. These claims predate the EPL and take precedence over it and are indicated by the orange rectangular perimeter in Figure 10. The claims expired on August 27, 2019, but are still active pending renewal.

Mr. Hoppe has also applied for two mining licenses overlapping a small portion of the northern portion of ML 200 totalling approximately 28.4 Ha (red rectangles on Figure 10). As at the effective date of this report, these applications had not been granted.

There are no other factors or risks known to the authors that might affect NMI's right or ability to perform work on ML 200.

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5. Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1. Accessibility

The town of Khorixas is connected to the capital city of Windhoek by approximately 450 km of paved road via Otjiwarongo and Outjo. Windhoek is the country's commercial and administrative centre and has international and regional airports with scheduled services to regional centres in southern Africa and Europe. Driving time from Windhoek to Khorixas is approximately 4.5 hours from Khorixas, the Lofdal project area can be reached via 25 km of secondary all weather gravel road. (Figure 11). Bush tracks provide good access to most parts of the project area and are generally negotiable by two-wheel drive (2WD) vehicles, although four-wheel drive (4WD) is occasionally required to cross gullies or wet areas during the rainy season.

Note: Coordinate system latitude and longitude coordinates in reference to the Bessel 1841 Spheroid, Central Meridian 17 degrees East
Source: MSA, (2024)
Figure 11 Location and road access to the Lofdal project area

5.2. Climate

North-western Namibia is an arid to semi-arid region. Rainfall is largely confined to the summer months (November to April) and averages 150 mm to 200 mm per annum. Average daytime high temperatures range from less than 25°C in June/July to more than 35°C from October to April and locally exceed 40°C during hot spells. Night-time lows reach 5°C in winter rising to about 20°C in summer. Sunshine averages more than 11

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hours per day during winter and eight to nine hours during the summer rainy season when there is frequent cloud cover. The climate is conducive to year-round operations.

The property lies within the catchment area of the Huab River and has little perennial surface water. The hydrological map of Namibia (van Wyck et al., undated) indicates that the project area is characterised by moderate to low water availability in the bedrock. Information from water boreholes in the area suggests that the water table is about 25 m below surface and comprises mainly fracture permeability in the crystalline basement rocks. Experience to date indicates that wells can supply sufficient water for the needs of exploration without compromising the requirements of local communities.

5.3. Local Resources and Infrastructure

Khorixas is a regional administrative centre. Local services include two fuel stations, hardware and general stores, a small supermarket and several convenience stores, a bank and facilities for basic vehicle servicing, welding and other trades. There is a small dirt airstrip but no scheduled air services. There are several tourist lodges in and near Khorixas offering accommodation, camping, and restaurants. NMI rents a 330 m² warehouse and a 7 600 m² surrounding yard supplied with municipal water and electricity that serves for equipment storage, local office, and drillhole core processing and storage (Figure 12). Core is processed at covered core logging areas at the tented field camp nearer the project area (Figure 13).

Source: Witley 2023 – top left, Witley (2020) – top right and bottom
Figure 12 Facilities in Khorixas for equipment storage, core logging and storage

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Source: Witley (2020)

Figure 13 Core processing facilities at the Lofdal Field Camp

The nearest large centres are Outjo, a town with a population of more than 6 000 that lies 120 km east of Khorixas; and Otjiwarongo, a town of approximately 20 000 people that lies 200 km east of Khorixas. Otjiwarongo is a regional commercial and service centre for, among other things, the former Okorusu fluorspar and Okanjande graphite mines.

There are local farms in and around the project area that raise cattle, sheep, goats, donkeys and horses. These farms have wells that can supply adequate water for exploration needs and together with Khorixas, provide a stable pool of workers that can be tapped for exploration requirements.

Khorixas is connected to Namibia's land telecommunications grid and has a Telecom Namibia office. Cellular telephone services are provided by MTC and internet data services are readily available through MTC as well.

Khorixas is connected to the national power grid via a 132 kV transmission line that runs to the east of Khorixas and northwards to the town of Kamanjab.

5.4. Physiography

The project area is characterised by low, gently rolling and sparsely vegetated hills with peaks ranging up to an altitude of approximately 1 030 m. There is an overall relief throughout the project area of slightly more than 100 m.

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Source: Swinden (2012)
Figure 14 Physiography of the project area showing typical low rolling hills and sparse vegetation

There is little soil, with much of the ground covered by residual gravel that closely reflects the composition of the underlying bedrock. This residuum is typically less than one metre thick on the high ground but thickens in the dry valleys. Outcrop is widespread throughout the area.

Vegetative cover includes a ubiquitous cover of native grass after the rainy season and numerous arid-adapted low shrubs. Wildlife is relatively sparse but includes springbok, kudu and gemsbok as well as baboons, elephants, zebras, leopards and various small mammals, lizards and snakes.

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6. History

The definitive geological mapping of the area was carried out by Frets (1969), who set out the general stratigraphic and tectonic framework. He identified and described the basement metamorphic rocks and the overlying sedimentary and volcanic rocks of the Damara Orogen and documented the alkaline intrusive plugs in the Lofdal area. Frets (1969) recognized and described the Oas quartz syenite and the silica-undersaturated syenitic rocks of the Lofdal complex. However, he did not recognize the associated carbonatites.

Other published accounts of carbonatites in the Lofdal area are found in Diehl (1990, 1992), Verwoerd (1993), and Woolley (2001) and accounts of the geological setting of carbonatites in the Lofdal area in Niku-Paavola et al. (2001) and Wall et al. (2008), which were summarised in Swinden and Siegfried (2011).

The current published geological map of the area is the 1:250 000 Fransfontein Sheet compiled in 2006 by the Geological Survey of Namibia (GSN) but it is too large a scale to be a useful base map at the detailed map scales needed for mineral exploration.

Historically, mineral exploration activities in the area have focused on copper, gold and tantalite associated with quartz veins and/or pegmatites hosted in the metasedimentary and metavolcanic gneisses of the Huab Complex. The copper and gold mineralization were generally interpreted to be related to faults and shears (GSN, 1992).

Small scale mining by way of shallow adits is evident in at least two locations within the ML immediately north of the project area (Figure 15). The adits were opened in the 1950's and one is reported to have mined mineralisation grading ten percent copper (GSN, 1992).

Source: Swinden (2012)
Figure 15 Adit at the former Lofdal Copper Mine with copper staining around the portal

Exploration for copper and gold in this area was conducted by Messina (Tvl) Development Co. Ltd. from 1974 to 1976 (Davidson, 1977). The GSN (1992) reported that diamond drilling in 1974 intersected cupriferous silicified zones, the best of which assayed 1.02%

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and 1.51% Cu over 1.95 m and 1.2 m respectively. Davidson (1977) interpreted the deposit to possibly represent the root zone of a largely eroded deposit. Tsumeb Corporation Ltd explored for base and precious metals between 1981 and 1986.

The area was prospected for gold by Anglo American Prospecting Services Namibia (Pty) Ltd. between 1987 and 1989. Reconnaissance bulk stream sediment samples were processed and analysed for gold content. Twenty-seven anomalous areas were selected for follow-up by detailed stream sediment sampling, soil sampling, rock sampling and geological mapping. This work did not yield any significant concentrations of gold and the project was terminated in 1989 (Marsh et. Al., 1989). The presence of copper mineralization, coupled with the presence of REE-bearing carbonatite dykes, abundant iron oxide mineralization, and magnetite-cemented diatreme breccias, led Lobo-Guerrero (2005) to suggest a potential for iron oxide copper gold (IOCG) type deposits in the area, which was influential in attracting the interest of NMI to the area. NMI explored for copper and gold in the area from 2005 to 2007 but did not delineate any IOCG targets from its regional exploration work. In 2008, NMI switched its focus to the potential for REE mineralization associated with the carbonatites.

Although not extensively described in the literature, carbonatite dykes have been reported in the area of Lofdal and Bergville farms at least since the early 1980's and were the focus of an exploration program for yttrium (Y) and rare earth elements (REE) by Rouna (Pty) Ltd. (Rouna) between 1981 and 1983. Following a reconnaissance radiometric survey and some rock sampling in 1981 and 1982 (Figure 16), attention was focused on anomalous responses in the area of farms Lofdal 491 and Bergville 491. The preliminary work by Rouna identified the presence of yttrium hosted in the mineral xenotime (YPO4) and demonstrated that radiometrics is an effective prospecting tool because of an association of rare earth elements with thorium. Detailed sampling in 1983 yielded ThO2 values ranging from 0.17% to 14.4% and yttrium from 207 ppm to 6,300 ppm with one analysis of 1.01% Y. One carbonatite dyke sample yielded anomalous contents of rare earths i.e., 0.82% Ce, 1.5% La and 0.74% Nd (Barbour, 1982). There are no other analyses of REE available from this phase of exploration.

Source: Swinden (2012)
Figure 16 Pit sampling of a carbonatite dyke from Rouna's exploration

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More recently, the Namibia Small Miners Assistance Centre held EPL 2821 over portions of Lofdal and Bergville farms for precious stones, semi-precious stones, precious metals, and base and rare metals from 2002 to 2004.

Mr. P. Siegfried investigated the greater Lofdal area in 1999 for Norsk Hydro ASA and in 2001 extensive sampling of the carbonatite dykes was carried out together with Dr. T. Mariano for the Canadian REE company Advanced Metals Research (AMR). REE mineralization and highly anomalous HREE were identified.

Geological investigations in the area by the GSN have been ongoing since V. Niku-Paavola began a Ph.D. research project on the carbonatites in the Lofdal area in 2004 at the Camborne School of Mines in the United Kingdom. This project was completed in 2014 (Do Cabo, 2014). Dr. R. Ellmies, previously of the GSN, has maintained an active research interest in this area and facilitated research particularly by students at the University of Namibia. Several B.Sc. theses describing aspects of the carbonatites and rare earth mineralization have been facilitated by the GSN (Ndalulilwa, 2009; Mutilifa, 2010; Shikongo, 2010). GSN's work in this area has provided significant new information on the unusual HREE enrichment at Lofdal and has yielded much detailed information about the mineralogy of the carbonatites and the related rare earth mineralization (Niku-Paavola et al., 2001: Wall et al., 2008; do Cabo and Ellmies, 2010).

In addition to the work at the University of Namibia, NMI has sponsored student work at a number of other universities. These include four B.Sc. (Honours) theses, respectively at Acadia University (Canada) (Kaul 2010), Dalhousie University (Canada) (O'Connor, 2011; Gaudet, 2012) and Stellenbosch University South Africa (Kruger, 2012). Two M.Sc. theses have been completed through the Camborne School of Mines (UK) (Loye, 2012, 2014). Two M.Sc. theses have been carried out on the petrology of the carbonatite intrusions, one at McGill University (Canada) (Bodeving, 2015) and a second at The University of St. Andrews (UK) (Robinson, 2020). A PhD study of the mineralized alteration zones and dykes was initiated at McGill University in 2013 but not completed. Preliminary results have been reported by Wollenberg et. al. (2016).

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7. Geological Setting and Mineralization

7.1. Regional Geology

The regional bedrock geology of north-western Namibia is defined by Archaean to Paleoproterozoic cratons to the north and south, the Congo and Kalahari cratons respectively, separated by a Neoproterozoic orogenic belt of Pan African affinity termed the Damara Fold Belt or Damara Orogen (Figure 17).

Note: Coordinate system is latitude and longitude in reference to the Bessel 1841 Spheroid, Central Meridian 17 degrees East
Source: Schneider, 2008
Figure 17 Cratons and orogenic belts in southern Africa

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The southern edge of the Congo Craton is exposed in three inliers in the Khorixas – Kamanjab area, termed respectively the Kamanjab, Braklaagte and Welwitschia inliers separated from each other by belts of younger volcanic and sedimentary rocks of the Damara Orogen.

The basement rocks in the Welwitschia Inlier were intruded post-tectonically by the Oas Syenite and the related Lofdal Carbonatite Complex, comprising syenite, nepheline syenite, phonolite and carbonatite.

Geological Survey of Namibia 2011

Note: Huab Metamorphic Complex Inliers (brown) are labelled: K : Kamanjab; B : Braklaage; W : Welwitschia.

Coordinate system is latitude and longitude in reference to the Bessel 1841 Spheroid, Central Meridian 17 degrees East

Source: (GSN, 2002)

Figure 18 General geology of Namibia

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7.2. Local Geology

The general geology of the Welwitschia Inlier, which hosts the Lofdal Carbonatite Complex, is shown in Figure 19. A more detailed geology map of the Lofdal area from recent geological mapping is shown in Figure 20. The principal geological units are briefly described below.

Source: (GSN, 2008)
Figure 19 General geology in the area of the Welwitschia Inlier

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Note: Detailed geology of Area 4 (black polygon) is shown in Figure 10 and detailed geology of Area 2B (red polygon) is shown in Figure 10.

Source: Swinden, 2014

Figure 20 Detailed geology of the area of the Lofdal Carbonatite complex

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7.2.1. The Huab Metamorphic Complex

The oldest rocks in the inlier are leucocratic granitic gneiss, banded paragneiss and quartzite, amphibolite, and mica/chlorite schist assigned to the Huab Metamorphic Complex (Frets, 1969) (Figure 21). Locally, mafic sills, dykes and stocks cut the sequence. The Huab Metamorphic Complex is polydeformed, affected by at least one phase of high-temperature isoclinal folding and is locally migmatised. The Huab Metamorphic Complex has not been directly dated but is considered to be about 2.0 billion years (Ga) old as it is intruded post-tectonically by the Fransfontein Granite, which has been imprecisely dated by U/Pb with two discordia lines giving ages of 1871 ±30 million years (Ma) and 1730 ±30 Ma (Burger et al., 1976).

Looking north across the Huab Metamorphic Complex in the Huab Welwitschia Inlier. High ground in distance is underlain by Damara Orogen sedimentary rocks.

Banded grey leucocratic granitic gneiss of the Huab Metamorphic Complex.

Source: Swinden, 2012

Figure 21 General topography and outcrop appearance of the Huab metamorphic Complex

7.2.2. The Damara Orogen

The Welwitschia Inlier is overlain to the north and south by sedimentary and volcanic rocks of the Damara Orogen. To the north it is in fault contact with clastic sedimentary rocks of the Mulden Group. To the south it is unconformably and/or structurally overlain by volcanic and sedimentary rocks of the Nosib and Swakop groups.

The Damara sequences represent a Pan African orogenic belt between the Congo and Kalahari Cratons. The basal successions (Nosib Group) comprise quartzite, arkose, conglomerate and subordinate calc-silicate and limestone that were laid down in, or marginal to, the intra-continental rifts. Locally, alkaline ignimbrite and associated subvolcanic intrusions are present (Naauwpoort Formation) the basal part of which has been dated by U-Pb single zircon as 752 ±7 Ma (de Kock et al., 2000). Early rift sedimentation was followed by widespread carbonate shelf and slope deposition, which grades laterally into deep water clastic sediments with local accumulations of within-plate basic volcanic rocks (i.e., the Swakop Group). Subsequent subduction and continental collision resulted in widespread deposition of molasse (Mulden Group).

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7.2.3. Early Damaran Alkaline / Carbonatitic Intrusions

At about 760 Ma, more or less contemporaneously with eruption of the early Damaran alkalic Naauwpoort Formation volcanic rocks, a suite of alkali silicate rocks and carbonatites were emplaced in the Huab Metamorphic Complex.

7.2.3.1. The Oas Syenite

The largest of these bodies is the Oas Syenite, first described by Frets (1969). It underlies approximately $20\mathrm{km}^2$ immediately south of the Lofdal project area, and comprises a dominantly coarse grained, alkali feldspar, sodium plagioclase, hornblende and quartz syenite (Figure 22). Apatite and sphene are important accessory minerals (Frets, 1969).

The Oas Syenite intrudes the basement gneisses and the basal sedimentary rocks of the Naauwpoort Formation (Frets, 1969) but is apparently overlain by Damaran limestones from higher in the sequence. The Oas Syenite has been dated by U/Pb in zircon as 756 $\pm 2$ Ma (Hoffman et al., 1996) and by U/Pb in titanite as $758 \pm 4$ Ma (Jung et al., 2007) and is therefore approximately coeval with the Naauwpoort Formation volcanic rocks.

Source: Swinden, 2012
Figure 22 Coarse grained Oas Syenite (alkali feldspar, amphibole and mica)

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7.2.3.2. The Lofdal Carbonatite Complex

Frets (1969) mapped a body of nepheline syenite in the southern part of farm Lofdal 491, intruding the gneiss complex. He noted a number of smaller satellite plugs, consisting dominantly of medium to coarse-grained leucocratic nepheline syenite as well as the prevalence of calcite and siderite in cracks and marginal facies of the intrusive.

More recent work has shown that the intrusive complex at Lofdal is more complicated than envisaged by Frets (1969) and comprises an assemblage of nepheline syenite and carbonatite intrusive plugs, dykes and hydrothermal alteration with related phonolite dykes and breccias defining an intrusive complex that appears to underlie an area of more than 200 km².

The regional setting of the Lofdal Carbonatite Complex is shown in Figure 20. The most important primary lithologies are nepheline syenite, phonolite, breccias and carbonatites as described below.

7.2.3.2.1. Nepheline Syenite

Nepheline syenites in the Lofdal area are medium to coarse grained, locally porphyritic syenites dominated by alkali feldspar, nepheline, sericite and biotite (O'Connor, 2011) (Figure 27). The original syenite intrusion mapped by Frets (1969) is, in fact, composite, comprising a carapace of syenite, which is intruded from below by carbonatite (the Main intrusion, see below). This results in a surface map pattern dominated by syenite. Syenite occurs in a number of satellite intrusions where it displays a wide variety of textures including very coarse-grained pegmatitic syenite phases. The syenite intrusions are locally cut by phonolite and carbonatite dykes, and fragments of the syenite are incorporated in the Lofdal breccias. The syenites are typically undeformed but locally exhibit mild shearing, characterised by development of a cleavage and alignment of feldspar phenocrysts. The Lofdal nepheline syenite has been dated by U/Pb in magmatic titanite as 754 ±8Ma (Jung et al., 2007) and it is therefore coeval with both the Oas Syenite and the Naauwpoort Formation volcanic rocks.

7.2.3.2.2. Phonolite

Phonolite (nepheline, alkali feldspar) dykes are widespread in the Lofdal area. They are dominantly northeast striking with fine-grained to moderately porphyritic with locally a trachytic texture (Figure 23). Where phenocrysts are present, they are dominantly alkali feldspar lathes up to two mm long with lesser nepheline. Phonolite dykes have been observed to cut syenite and breccia and they are typically closely associated with carbonatite dykes throughout the area. A comparative petrographic and mineralogical study of the syenites and phonolites led O'Connor (2011) to conclude that they are likely co-magmatic.

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Source: Swinden, 2012
Figure 23 Examples of nepheline syenite and phonolite dyke

Slightly porphyritic phonolite dyke. Phenocrysts are alkali feldspar and nepheline.

7.2.3.2.3. Lofdal Breccias

Frets (1969) first noted the presence of a very coarse breccia associated with the nepheline syenites, which he described as "closely packed angular fragments of gneiss which vary in size between one cm and 50 cm, embedded in a fine-grained, contaminated facies of the syenite". He suggested that it indicated a forceful intrusion of the syenite.

Geological mapping in the area has demonstrated that these breccias are widespread and are associated with virtually all the syenite intrusions identified. Although locally dominated by country rock fragments as described by Frets (1969), in other areas they are dominated by syenite fragments. The breccias range from polylithic (basement clasts) to monolithic (syenite clasts) and are typically unsorted, angular, and chaotic (Figure 24). Locally, the breccias are intruded by carbonatite and phonolite. They clearly post-date the intrusion of the syenite plugs, as the syenites form fragments in them, but they must be closely related in time as phonolite dykes are locally observed to cut the breccias. There are no carbonatite fragments in the breccias and they apparently predate the carbonatite intrusion.

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Coarse grained, clast-supported, unsorted syenite breccia near the Emanya intrusion with a small carbonatite dyke under the hammer
Figure 24 Examples of Lofdal breccias

Polymictic breccia exposed in a quarry near the Main intrusion

Source: Swinden, 2012

7.2.3.2.4. Carbonatites

Carbonatite dykes have been reported in the Lofdal area for a considerable time (Barbour, 1982; Verwoerd, 1993; Miller, 2008). However, the recognition of larger, plug-like carbonatite intrusions in the area is relatively recent. The composite syenite- carbonatite plug that is now referred to as the "Main" intrusion appears on a map by Barbour (1982) but its full extent and significance was only first recognized by the GSN geologists (V. Do Cabo, pers. comm.). A second, smaller intrusion about 4.5 km to the southwest, now referred to as the "Emanya" intrusion or plug, was only discovered in 2008 (Figure 25). Regional geological investigations between 2008 and 2010 have shown that there are literally hundreds of carbonatite dykes and fentitised-carbonatised alteration structures over an area of about 200 km². Wall, et al., (2008) dated xenotime in the carbonatites by U-Pb and obtained an age of 765 ±16 Ma, indicating that they are approximately coeval with the Oas Syenite and the Lofdal nepheline syenites.

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White sovite of the Main Intrusion intrudes a carapace of nepheline syenite (ledge at top of outcrop).

Dark reddish-brown carbonatite is characteristic of the Emanya intrusion.

Source: Swinden, 2012

Figure 25 Examples of carbonatite from the Main and Emanya intrusions

The Main intrusion is the largest carbonatite body found to date in the complex. Its outcrop area forms an ovoid with long axis of about two km and an area of about 1.5 km²

Figure 20). The Main intrusion carbonatites are dominantly coarse-grained white sovite consisting mainly of calcite with lesser aegerine, apatite and magnetite, and trace feldspars, sulphide minerals and pyrochlore (Gaudet, 2013; Bodeving, 2015). The carbonatite intrudes nepheline syenite and Lofdal breccias, which form an outcrop carapace on top of the carbonatite at the present level of exposure (Figure 25). Because of this, syenite, and to a lesser extent breccia, dominate the outcrop pattern in the area of the intrusion and it is not surprising that previous workers mapped this body as dominantly syenite (e.g., Frets, 1969). The Main intrusion is relatively uniform geochemically, with low iron contents and exhibits a LREE-enriched rare earth element distribution that is typical of carbonatite magmas but too low in absolute REE concentration to be of economic interest.

The Emanya intrusion is located about 3.8 km southwest of the Main intrusion (Figure 20). It comprises a main body, roughly circular in outcrop with a diameter of approximately 350 m, as well as several smaller satellite bodies within a 450 m radius. The carbonatites in this intrusion are calcitic but contrast with the Main intrusion in that they are finer grained and dominantly brown to reddish-brown on outcrop surfaces with abundant iron oxide throughout (Kruger, 2012) (Figure 25). Fluorite is locally present in veinlets. On average, the Emanya carbonatites contain approximately 8.6 times more LREE and 3.6 times more HREE than the Main intrusion. The REE in Emanya are fractionated in favour of the LREE compared to the Main intrusion.

Carbonatite "dykes" have been mapped over an area of more than 200 km² throughout the Lofdal Carbonatite Complex. Although typically mapped and referred to as "dykes", these appear to be dominantly hydrothermal and/or carbothermal vein systems, resulting from fluid expulsion from the magmas during crystallisation. They typically follow the structural grain of the country rocks, striking in a northeasterly direction and dipping steeply to the south. They exhibit a wide range of lithological and alteration characteristics, ranging from <10 cm to several 10's of metres wide and have a wide range of colour variation on weathered surfaces, from white and grey, through shades of brown, red and yellow (Figure 26). They are closely associated with phonolite dykes throughout the area, often occurring

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together in the same structure (with the 'dykes' always the later phase) or in closely spaced parallel dyke swarms.

There is considerable variability in the internal structure of the "dykes". Many are uniform in character, exhibiting little internal banding or colour variation. Most however, exhibit some internal structure, commonly colour- and/or compositional-banding on a scale ranging from millimetres to centimetres.

The alteration in these "dyke" systems is variable but locally intense, characterised by an early pervasive albitisation follow by brittle fracturing and infusion of carbonate minerals and micas. HREE mineralization is associated with the later stages of alteration. In some areas the alteration systems are 10's of metres wide.

A 10 m wide massive brown carbonatite dyke

Laminated red and yellow carbonatite dyke with a band of albitite along the right side (under the hammer head).

Source: Swinden, 2012

Figure 26 Examples of brown and red to yellow carbonatite dykes

7.3. Structural Setting

Rocks of the Huab Metamorphic Complex were polydeformed and metamorphosed prior to intrusion of the Fransfontein Granite at about 1.7 Ga. They were subsequently affected by extensional tectonics during the rifting event that initiated the Damara Orogen at 850 Ma to 750 Ma and transpression during the orogenic events that accompanied the Damara Orogeny from 580 Ma to 500 Ma. The effects of both Neoproterozoic rifting and early Paleozoic transpression are recorded in the Lofdal carbonatites.

A comprehensive structural interpretation of the area was undertaken by NPA Fugro (2010) on behalf of NMI, integrating hyperspectral and Landsat data with all available airborne geophysical surveys as well as NMI geological and geochemical databases to produce a high level interpretation of structural features in the Welwitschia Inlier (Figure 27). This interpretation identifies regional structures in the basement that probably reflect the pre-Damara history of these rocks. It also identifies a series of sinuous NE-SW striking major fault structures that systematically offset the basement structures in a sinistral sense. These structures are locally offset by a series of NNE-SSW striking dextral structures and NW-SE striking structures.

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Note: Base is hyperspectral image. Main and Emanya intrusions and carbonatite dykes are shown in white.
Source: NPA-Fugro, 2010
Figure 27 Structural elements of the Lofdal area, interpreted from Landsat and hyperspectral data

The intrusive complex at Lofdal shows a close relationship to the structural grain of the basement. The Main and Emanya intrusions are located between large sinistral structures and the phonolite and carbonatite dykes and veins in the complex are typically structurally aligned with these ENE- and NNE-trending structures (Figure 27). The intrusive complex seems, therefore, to have exploited regional structures during emplacement and subsequent hydrothermal alteration activity.

Structures within the carbonatites appear to record both extensional and transpressional events. The repeated injection of first phonolite and then several phases of carbonatite, and hydrothermal fluids into many of the structures may be the result of repeated opening of pre-existing structures during the extensional regime that prevailed at about 750 Ma. However, there is also evidence of transpressional deformation within the dykes, with minor structures (folds, shear bands) indicating transpression with the same sense of shear as is interpreted for the major structures. It may be that at least some of the deformation recorded in the dykes is Damara in age. It seems likely that the intrusions and the associated hydrothermal alteration were focused by extension faults that determined the location and orientation of the intrusions and provided pathways for the escaping hydrothermal fluids. These faults and fractures may have been reactivated during the Damara Orogeny, producing the observed structures.

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7.4. REE Mineralization

7.4.1. Regional Setting

Exploration at Lofdal has demonstrated that there is widespread REE mineralization related to intrusion of the Lofdal Carbonatite Complex and that many occurrences are specifically enriched in HREE. The regional setting of REE mineralization was evaluated through an extensive regional surface grab sampling program (documented in Swinden and Siegfried, 2011), geological mapping with locally detailed lithogeochemical sampling, and core drilling. Figure 28 illustrates the distribution of anomalous concentrations of TREE+Y in 3,764 outcrop grab samples collected between 2008 and 2011. The REE mineralization at Lofdal occurs mainly within a NE-SW trending corridor approximately 20 km long and 5 km wide, the axis of which is occupied by the Main and Emanya intrusions. Mineralization occurs at a district scale over an area of at least 200 km².

Figure 28 Distribution of lithogeochemical grab samples in the Lofdal area

Note: Color-coded for total REE + Y. Main intrusion is large grey body. Emanya plug (small grey body) to the southwest
Source: Landsat Geocover Mosaic, 2000

There is considerable variation in both the absolute concentrations of REE, and in the relative proportions of LREE versus HREE in mineralised samples. As a general rule, the Main intrusion shows a typical carbonatite REE profile of LREE enrichment but has very low overall concentrations of all REE.

The Emanya intrusion shows a trend towards higher TREE+Y, which is mainly a result of LREE-enrichment, with little enrichment in HREE. The dykes and related alteration lithologies show a wide variation with much stronger enrichment trends in both LREE and HREE (Swinden and Siegfried, 2011).

Studies of both surface outcrops and drillhole cores have demonstrated that the HREE-rich mineralization is not principally hosted by carbonatites, but typically occurs in hydrothermal and/or carbothermal alteration zones that are localized in key basement structures irrespective of host lithologies (do Cabo et al., 2011; Swinden and Burton, 2012;

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Wollenberg et 2016). The delineation core drilling for mineral resource estimation in Areas 2B and 4 shows that the HREE-rich mineralization is locally continuous in three dimensions over significant strike and dip extents.

Despite the complexity of the REE distribution overall, there is a regularity to the distribution of the most HREE-enriched samples. Figure 29 shows the enrichment of heavy rare earths as $(\mathrm{HREE} + \mathrm{Y}) / (\mathrm{TREE} + \mathrm{Y})$ expressed as percentage, irrespective of grade.

Note: Most HREE-enriched samples plot along linear trends that are interpreted to reflect structures that provided fluid pathways during the hydrothermal event. Main and Emanya intrusions are shown in light brown; alteration intensity in shades of grey (see also Figure 31 and Figure 33). Black lines are structures interpreted from remote sensing and geophysical data.

Source: Swinden, 2011

Figure 29 Lithogeochemical grab samples plotted on the basis of $(\mathrm{HREE} + \mathrm{Y}) / (\mathrm{TREE} + \mathrm{Y})$

Figure 29 demonstrates a number of important distribution characteristics of the HREE-enriched mineralization:

The most HREE-enriched samples tend to be concentrated in linear belts, which very often coincide with the traces of fault structures interpreted from remote sensing data. These are interpreted to be the structures that provided fluid pathways or conduits for the HREE-rich hydrothermal fluids.
Even where overall grades are low in these structures, the HREE enrichment remains high, emphasizing the potential of these zones for concentrations of HREE-rich mineralization.
The drill targets in Areas 2B, 4 and 5 stand out as zones of HREE-enrichment.

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7.4.2. Mineralization in Area 4

The location of Area 4 in the Lofdal Carbonatite Complex is shown in Figure 20 and a detailed geological map of Area 4 is shown in Figure 30 and Figure 31.

The Huab Metamorphic Complex in this area is dominated by quartzo-feldspathic gneiss and metasedimentary grey gneiss with lesser amphibolite and pegmatite. The gneisses strike approximately ENE-WSW and generally dip steeply southwards.

The geological element of principal economic interest in Area 4 is a major fault, first interpreted from remote sensing data (NPA-Fugro, 2010), that strikes ENE-WSW and dips to the south, bisecting the area. The fault can be traced for several kilometres east and west of Area 4 and the offset of geological elements interpreted from hyperspectral data indicates that it has a sinistral sense of movement. At surface, the fault system is marked by carbonate veining and extensive hydrothermal alteration, dominantly albitisation and carbonatization accompanied by biotite ± phlogopite and iron oxides. Mapping the alteration intensity associated with the structure shows that there is a core of intense alteration, within which rocks have been completely converted to albitite and are cut by carbonatite and highly carbonic alteration. In this intense alteration zone, all original textures have been destroyed and crackle breccias with altered, albitised clasts set in a matrix of carbonatite and/or iron oxides are common. Surrounding this intensely altered core is a halo of less intense alteration, in which the rocks are bleached and albitised, but retain some original textures.

The outline of the alteration zone is highly irregular at map scale (Figure 30). The intense alteration in the core is typically between $15\mathrm{m}$ and $30\mathrm{m}$ wide on the surface (not true width). The less intense alteration halo exhibits gradational and diffuse contacts with the wall rocks and is typically on the order of $50\mathrm{m}$ to $60\mathrm{m}$ wide at surface but can range to more than $100\mathrm{m}$ wide. At the eastern side of Area 4, the fault zone bifurcates, with the main system branching in a slightly more northerly direction and a splay of the fault continuing in an ENE direction. There is alteration and mineralization associated with both splays, but the southern splay appears to decrease in intensity along strike and alteration appears to die out within a few hundred metres.

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Note: The mineralised structure reports high values in both the HREE (represented by Dy in ppm) and in percentage of HREE in TREE
Source: Swinden, 2021
Figure 30 Geology of Area 4 with Dysprosium (Dy) grade in surface grab samples

A large number of carbonatite veins and carbonatic alteration zones have been mapped in Area 4. Although strike directions are dominantly NE-SW and NNE-SSW following the dominant structural grain of the basement, other directions are locally seen. Outside the central alteration zone, carbonatite veins are thin, (<1 m wide) and do not exhibit significant alteration beyond their margins. However, within the alteration zone, they are more continuous and alteration is ubiquitous.

A more or less continuous zone of albite-carbonate alteration with significant grades of REE has been traced by mapping, trenching and drilling for more than 1,100 m along strike and regional geological mapping to the east and west indicates that it continues for several kilometres beyond Area 4. Within this zone, the intense alteration typically thickens and thins, and locally forms lensoid bodies that can extend on surface up to about 100 m long and 10 m wide.

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The mineralization in Area 4 occupies a structurally-controlled, linear alteration zone. The Area 4 alteration zone is the largest and best mineralised and is clearly manifested and easily mappable in surface outcrops by variably intense albitisation and brown carbonization with locally abundant phlogopite. Grab samples from outcrops typically return highly anomalous values of HREE (Figure 30) and also have a very high HREE/TREE ratio (Figure 31).

Note: The mineralised structure returns high values in both the HREE and in percentage of HREE
Source: Swinden, 2021
Figure 31 (HREE+Y)/(TREE+Y)% in surface grab samples in Area 4

7.4.3. Mineralization in Area 2B

The location of Area 2B in the Lofdal Carbonatite Complex is shown in Figure 20 and a detailed geological map of this area is shown in Figure 32 and Figure 33.

The Huab Metamorphic Complex in Area 2 is dominated by amphibolitic schist interbanded at outcrop scale with leucocratic quartzo-feldspathic paragneiss and muscovite schist and locally intruded by coarse grained granitic pegmatites. The rocks are complexly folded on a fine scale.

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Figure 32 Geology of Area 2B with dysprosium (Dy) grade in surface grab samples

Note: Alteration related to the Area 2B mineralization is shown in shades of grey
Source: Swinden, 2021

Phonolite and carbonatite dykes and related alteration zones of variable orientation and thickness are common in the area. Carbonatite dykes average a few cm in width but carbonatitic and albititic alteration zones can range up to more than 10 metres in width in outcrop.

Dykes and alteration zones in this area dominantly trend from NE-SW to NNE-SSW and generally are at a considerable angle to the structural grain of the basement, which in this area trends from E-W to ESE-WNW. The area is bounded to the north and south by major sinistral faults interpreted from remote sensing data (Fugro, 2010) and it may be that the dominantly NE-trends of dykes and alteration zones in Area 2B reflect fracture systems related to these linked faults.

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Source: Swinden, 2021
Figure 33 (HREE+Y)/(TREE+Y) % in surface grab samples in Area 2

The principal mineralization in Area 2 is the 2B zone (Figure 32), a wide zone of hydrothermal alteration and carbonatite intrusion. Like Area 4, the mineralization is characteristically enriched in HREE and samples throughout the alteration zone show a very high ratio of HREE to TREE (Figure 33). The mineralized zone has been traced in outcrop along a strike length of more than $600\mathrm{m}$ and remote sensing information and regional sampling results suggest that the zone may ultimately have a strike length of more than $3\mathrm{km}$ . The width of the zone in outcrop is variable. At its southern end, the width of the zone of alteration and carbonatization ranges from about $20\mathrm{m}$ to $35\mathrm{m}$ but thins to less than $10\mathrm{m}$ in the central section where it bifurcates into two separate zones. At the northern end, where the zone of alteration and carbonatization is again amalgamated, it is more than $60\mathrm{m}$ wide. In outcrop, it comprises a zone of massive carbonatite dykes, within a complex envelope of hydrothermal alteration and brecciation. The southern part of the zone is dominantly brown carbonatite and related alteration. Alteration zone lithologies include massive albitite, localized zones of green phlogopitic fenite, brown stained albitite breccias infused with carbonatite, as well as altered mafic schist that has been carbonatized, variably albitized, and intruded by carbonate veinlets. Although massive and continuous across strike at its southern end, the zone bifurcates at surface in its central

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section with a hanging wall zone of dominantly carbonatite and related alteration and a footwall zone that includes considerable altered, carbonatized schist and carbonatized stockwork.

The REE mineralization in the Area 2B zone is restricted to the zone of alteration and carbonatization. In drillhole core, the zone is seen to consist of a zone of intense brecciation and hydrothermal alteration. The best assay values are related to late veining and alteration that cuts most of the pre-existing alteration lithologies.

Shearing is very common in the alteration zones. In most sections, there is a prominent shear zone at or near the footwall, which is itself variably albitized and carbonatized. The shear zones range from centimetres to as much as five metres wide and the shear fabrics are cut by both albitite and carbonatite suggesting that they represent structures that predate the mineralization. There is a main footwall shear zone in most sections, which may be the controlling structure for much of the alteration and mineralization.

7.4.4. Nature of the Alteration

The alteration in Areas 4 and 2B is geologically and mineralogically very similar. In drillhole core, the zones are characterised visually by the bleaching and reddening that accompanies the alteration (Figure 34). The boundaries of the alteration zones are locally sheared and/or intensely broken, clearly indicating the structural nature of the zones. The structure is internally complex with multiple internal shears that are typically micaceous, characterised by the development of black biotite, phlogopite and chlorite as well as calcite.

The alteration and contained mineralization are characterised by both radiometric and geochemical anomalies. The presence of Th results in a generally elevated radiometric signature in the alteration zones, although the Th is not always spatially associated with the REE. Geochemically, the alteration zones are characterised by elevated concentrations of the HREE, Y and $\mathrm{P}_2\mathrm{O}_5$ as well as Nb and Zr, although, as with Th, the Nb and Zr concentrations do not closely correlate with the HREE on a sample-by-sample basis (Figure 35). There are no correlations between the HREE and the LREE. Visual radiometric and geochemical characteristics allow the alteration zone to be readily traced between drillholes both down-section and along strike (Figure 35).

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Source: Swinden, 2012

Figure 34 Colour anomaly in drillhole core associated with the Main zone alteration in Area 4

Note: Element concentrations in ppm except $P_2O_5$ in weight %. Depth in metres (vertical axis). Alteration zone is defined visually in drillhole core, Main zone mineralization by assays correlated with visual alteration.
Source: Swinden, 2014
Figure 35 Schematic illustration of geochemical and radiometric anomalies associated with the Area 4 alteration zone in drillhole NLOFDH4047

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Figure 36 Schematic cross section of the upper part of the Area 4 Main alteration zone from drillhole data

Note: Overlapping drillholes are NLOFDH4016 and 4037 which were twinned. Colour bars along drillholes represent concentrations of Y ranging from highest (red) through orange, pink, yellow and green to lowest (grey). Heavy dashed lines are structural boundaries of the Main zone.
Source: Swinden, 2012

Rocks within the alteration zones are typically completely replaced by an assemblage containing variable proportions of albite, quartz, biotite/phlogopite, chlorite, calcite and iron oxides. The alteration is characterised by a pervasive background albitisation which has converted the entire rock mass to fine grained albitite. The early albitite has been brecciated and overprinted by a second generation of albite, calcite, brown calcite and dolomite, which locally form thin brown carbonatite veins, red iron oxides, and green to black biotite/phlogopite and chlorite (Figure 37). The alteration is typically texture-destructive and pervasive.

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Note:
1. Early albite (white) brecciated and infilled by hydrothermal carbonate (brown)
2. Alteration in drillhole core. White albitization overprinted by brown carbonate/iron oxides
3. Narrow highly mineralized iron oxide-rich vein cutting albitized, mica rich alteration
Source: Swinden, 2010

Figure 37 Typical alteration and mineralization in Area 2B

The alteration halo is typically broader than the mineralised intervals. Alteration relationships are complex, but there are regularities and consistencies to the distribution of mineralization. Mineralization is variously found in red to pink albite-rich veins and patches, black mica- and chlorite-rich alteration veins and shear zones, tan and silver-grey variegated albitite, white to grey dolomite, white, grey or yellow albitite, or brown carbonatite (Figure 38). Typically, several styles of mineralised alteration occur within the same drillhole and where the alteration is most complex; there are clear overprinting relationships between different generations of alteration. The main unifying geological characteristic of the higher-grade mineralization is the presence of complex overprinting relationships, reflecting multiple alteration events. Almost invariably, when paragenetic relationships between different generations of alteration are observed, the alteration associated with highest-grade mineralization is the latest event.

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Note:
Figure 38 Examples of Area 4 alteration in drillhole core

A: Dark grey albite-calcite-biotite/phlogopite matrix with tan xenotime-rich alteration;
B: Red vein network rich in HREE cuts a brecciated white albitite;
C: HREE-rich red alteration cross-cuts shear fabric in a mica-rich shear zone

Source: Swinden, 2012

In areas where significant grades and widths of HREE mineralization are intersected by drilling, the mineralised zones typically exhibit an increase in the background HREE content of more than threefold over the background in unmineralised rocks. These intervals generally contain local areas of high-grade mineralization that raise the overall grade from anomalous to economically- significant. High-grade mineralization may take various forms, the most common of which are veinlets, vein networks, alteration patches and micro-breccia veins (Figure 39). The higher-grade mineralization is characterised by extreme HREE enrichment ([HREE+Y]/[TREE+Y] >90%) and this enrichment typically extends beyond the boundaries of the higher grades into adjacent rocks where the overall grades are lower.

Note:
Figure 39 High grade mineralization in Area 4 alteration in drillhole core

A: Red alteration patch in centre contains >10% Y;
B: micro-breccia vein trends from lower left to upper right and matrix is rich in HREE

Source: Swinden, 2012

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7.4.5. Mineralogy

Petrographic and scanning electron microscope (SEM) observations provide detailed information on the mineralogy of the mineralized zones. Most of the detailed work was carried out on Area 4, but regional studies in other alteration zones indicate that the mineralogical characteristics are similar throughout the property.

Detailed mineralogical studies of the Lofdal carbonatites and associated mineralization were first carried out by Mariano (2001) who studied a mineralised sample from a carbonatite dyke. His petrographic and cathodoluminescence studies showed the principal REE-bearing mineral to be xenotime, with associated monazite, parisite, apatite and thorite.

A petrographic study of 18 variably mineralised samples from outcrops in the Lofdal area was carried out by Schandl (2010). The principal HREE mineral identified was xenotime with minor aeschynite (Y). Minor amounts of LREE minerals were identified including bastnaesite, parisite, synchysite, monazite and a single occurrence of allanite. Schandl (2010) noted that the low Th content of most REE minerals may signal a hydrothermal origin and identified secondary albite, riebeckite and aegerine which were interpreted as evidence of sodic fenitisation.

Detailed mineralogical studies on a suite of outcrop samples by V. do Cabo of the GSN, including whole rock geochemistry and scanning electron microprobe studies, confirmed the dominance of xenotime in mineralised samples and identified a suite of accessory minerals that include zircon, monazite-(Ce), synchysite-(Ce), thorite, apatite and rutile in a calcite-albite-quartz-chlorite-Fe-oxide gangue (Wall et al., 2008).

Detailed mineralogical studies of mineralised drillhole core from Area 4 were undertaken by Dr. James Clark of Applied Petrographics (Clark, 2012). Petrographic and SEM studies show that the gangue to the mineralization comprises mainly albite, phlogopite/biotite and chlorite, calcite iron oxides and quartz. In most mineralised sections, the background lithology is albitite, dominated by coarse to fine crystals and crystal fragments of albite. The textures in the albitite indicate that it has been granulated resulting in remnant coarse albite crystals and crystal aggregates within a comminuted matrix of rock flour flooded by calcite and iron oxide (Figure 40).

Note: Aggregates of xenotime and zircon (high relief, examples highlighted), in association with calcite, distributed along micaceous stringers and intergranular to granulated albite. The dark brown phase near photo centre is niobian rutile. FOV=1.35 mm. Cross polars on left, plane light on right.
Source: Clark, 2012
Figure 40 Albitite with aggregates of xenotime and zircon

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Accessory minerals, including REE-bearing phases, occur in the matrix, in hairline fractures and along shear fabrics defined by biotite/phlogopite. The working hypothesis is that pervasive albitisation was an early alteration event, and that the albitites were affected by further movement on the host structures, resulting in brecciation and the introduction of new hydrothermal fluids which resulted in overprinting alteration and mineralization.

The most abundant accessory minerals are Nb-rutile and rutile, zircon, thorite and apatite. Minor amounts of pyrite, ilmenite and galena were also identified.

The most abundant REE-bearing mineral is xenotime, which occurs in more than 80% of the samples examined from Area 4. Synchysite-(Ce) and Synchysite-(Y) are common, although minor phases, and minor amounts of monazite-(Ce), bastnaesite-(Ce), parisite-(Ce) and aeschynite-(Y) are locally present as well as a number of REE- bearing phases that have not yet been identified. SEM spectra suggest that some REE may be present in zircon and thorite.

SEM backscatter images show that the mineralization is typically fine grained. Individual grains and grain aggregates are locally >100 µm but are typically less than 50 µm. Xenotime is locally intergrown with zircon, rutile and thorite on a fine scale (Figure 41).

Note:

A: Xenotime crystals and aggregates in a lens of calcite and phlogopite/chlorite. Rutile and niobian rutile are present, along with minor accessory apatite and iron oxide after sulphide. Scale bar is 200 µm;
B: Xenotime aggregates locally envelop earlier-crystallising zircon. Minor synchysite is intergranular to albite and biotite, and locally in edge contact with xenotime. Scale bar is 50 µm;
C: Xenotime deposition along a rounded edge of the zircon crystal at photo centre. Xenotime appears to nucleate on the earlier zircon. Phlogopite and granulated albite are the gangue phases. Scale bar is 20 µm;
D: Monophase xenotime aggregates. Scale bar is 200 µm.

Source: Clark, 2012

Figure 41 Backscatter images of Area 4 mineralization

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Loye (2014) carried out a detailed study of the Area 4 alteration and mineralization using observations of drillhole core, geochemistry, cathodoluminescence, SEM and microprobe data. He recognized six different modes of occurrence of xenotime and ascribed these to an extended process of HREE mineralization and remobilisation spanning the late magmatic and hydrothermal phases of the intrusions. His model involves early ground preparation of key structures by alkalic fluids expelled during crystallisation of the nepheline syenites, which resulted in pervasive and widespread albitisation. Continued movement on these structures resulted in brecciation of the brittle albitites. Fluids exsolved from carbonatite magmas utilised the fluid pathways created by the brecciation, overprinted the albitites with a complex alteration assemblage that included dolomite and ankerite, biotite/phlogopite, iron oxides and pyrite, and a variety of accessory phases, and introduced HREE rich mineralization. Early alteration was dominated by calcite and dolomite, and late alteration by ankerite. Late alteration fluids re-worked the early alteration assemblages remobilizing and redistributing previously present REE.

7.4.6. Thorium

The presence of thorium (Th) can potentially be problematic in carbonatite-associated REE deposits because of its radioactive nature. Alteration associated with REE mineralization in the Lofdal Carbonatite Complex is variably anomalous in Th, and this largely contributes to the regional Th airborne radiometric anomaly that defines the area of interest. It also provides a convenient and important prospecting and evaluation tool on the ground, as most carbonatites and their associated alteration have elevated radiometric signatures.

The alteration zones in Areas 4 and 2B typically carry anomalous concentrations of Th (approximately 2% of drill samples returned >1,000 ppm Th), and the zones give a low-level radiometric response which is generally a good guide to mineralised alteration zones. However, overall, there is not a close geochemical relationship between the HREE and Th. Figure 42 shows the total HREE versus Th results for 5,940 drillhole core and trench samples from Area 4 and Area 2B.

Figure 42 THREE versus Th in trench and drillhole core samples

Note: Although Th is locally present in mineralised samples and radiometrics often provide a good guide to mineralization, there is no clear geochemical correlation between Th and the HREE.
Blue dots- Area 4; orange dots-Area 2B
Source: Swinden, 2021

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7.4.7. Mineralization Summary

Mineralization in Areas 4 and 2B is structurally controlled and hydrothermal in origin. The host structures are first- and perhaps second-order basement structures that were apparently reactivated more than once during the mineralizing event. Repeated movement promoted the introduction of several generations of hydrothermal fluids, which resulted in a complex series of overprinting alteration events. The mineralization is dominantly present in xenotime and is interpreted to be related to the waning stages of hydrothermal alteration related to carbonatite intrusion. The highest-grade mineralization does not occupy a consistent position within the structural zones. It is interpreted to occupy structures within the zone that were still open during the last phases of hydrothermal alteration. The mineralised structures can be traced from hole to hole and are variably mineralised.

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8. Deposit Types

8.1. General Models for REE Mineralization in Carbonatites

Carbonatites and related, often undersaturated, silicate rocks originate in the earth's mantle through very low degrees of partial melting. They typically display geochemical enrichments in Ba, Nb, P, Fe, Ti, REE, F, Sr, Ta, Th, U and Zr. Carbonatites are important for a variety of economic mineral deposit types including REE (e.g., Bayan Obo, Mountain Pass), Nb (e.g., Araxa, Oka), Ti (e.g., Tapira), P (e.g., Araxa, Palabora), vermiculite, and fluorite (e.g., Okoruso). Carbonatite-associated deposits, including the giant Bayan Obo deposit in China, are the principal source of REE. Carbonatites are the focus of much current exploration for REE throughout the world.

A widely cited general model for the intrusion of a carbonatite complex (Le Bas, 1987) is shown in Figure 43. Although most complexes differ from each other in detail, this model provides a useful framework for description of observed mineralization at Lofdal. At Lofdal, the early silicate intrusions are dominantly syenite and nepheline syenite, rather than ijolite and urtite. Like the model, they are succeeded by a sovite intrusion (the Main intrusion) and by one and potentially more, subsidiary intrusive plugs, represented by the Emanya intrusion.

Similar to the model, Lofdal has abundant later stage dykes of both silicate (phonolite) and carbonatite, ranging in composition from calcitic through dolomitic and ankeritic phases. There is abundant fenitisation related to hydrothermal alteration around the margins of the intrusions and in basement structures that have served as pathways for both phonolite and carbonatite magmas and later hydrothermal fluids.

Note: Early silicate intrusions are intruded by carbonatite intrusions, which are cut by later carbonatitic dyke complexes.
Figure 43 General cross-sectional model for an alkali silicate-carbonate intrusive complex (after Le Bas, 1987)

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REE, particularly the LREE, are typically significantly enriched in carbonatites over normal crustal abundances – a result of their partial melting history in the mantle, subsequent concentration through fractional crystallisation in the crust and sub-solidus hydrothermal activity accompanying the intrusion. Locally, the enrichment in REE produces deposits of economic proportions. Mariano (1989) identified three types of mineralization that might be expected in a carbonatite complex:

Magmatic primary crystallisation of REE minerals;
Hydrothermal concentration; and
Supergene concentration. The rocks at Lofdal are not deeply weathered and there is a very low likelihood of extensive supergene enrichment. However, there is significant potential for both magmatic and hydrothermal deposits.

There appear to be at least two styles of mineralization on the Lofdal property:

Early LREE-enrichment in magmatic carbonatites, particularly the Main and Emanya intrusions: Absolute abundances in the Main intrusion do not appear to attain economically interesting grades but overall grades are higher in the Emanya intrusion. Associated sovite dykes are also significantly LREE-enriched; and
Late hydrothermal mineralization characterised by extreme HREE enrichment: This mineralization is dominantly structurally controlled and occupies hydrothermal alteration zones within major structures. This mineralization is characteristic of Area 4 and Area 2B where diamond drilling has outlined mineral resources.

8.2. Magmatic Mineralization

It is rare to find a REE deposit that has formed through primary crystallisation from carbonatite magma. The best-known example is the Mountain Pass deposit in California where a 1.4 Ga intrusive complex consisting of a total of eight plugs, ranging in composition from shonkinite to carbonatite, intrude Precambrian basement metamorphic rocks (Castor, 2008). The deposit reportedly has "current reserves" of more than 20 million tonnes of ore with an average grade of 8.9% rare-earth oxides (Castor, 2008). The ore typically contains 10% to 15% bastnäsite-(Ce) with subsidiary monazite and apatite, and is mostly composed of calcite, dolomite and barite, with generally minor amounts of other minerals. Texture and mineral paragenesis shows that bastnaesite and parasite are primary magmatic minerals.

The Mountain Pass carbonatite plug provides an analogue for potential LREE targets at Lofdal. There is widespread LREE-rich mineralization in both intrusive plugs and carbonatite dykes, particularly in the central intrusive core of the Lofdal complex. This mineralization has not been extensively explored to date, with the exception of seven drillholes testing the Emanya intrusion. The carbonatite intrusion at Mountain Pass is comparable in size to the Emanya intrusion and there is potential for the discovery of additional plugs of similar size at Lofdal. There continues to be significant potential for LREE-rich mineralization associated with the intrusion of the Lofdal carbonatites but to date, exploration has not focused on the LREE targets.

8.3. Hydrothermal Mineralization

There is abundant evidence, both observational and theoretical, that REE minerals are precipitated from hydrothermal solutions (Williams-Jones et al., 2013) and, according to Mariano (1989), this is the origin of REE minerals in most carbonatites. REE mineralization in carbonatites may result from the breakdown of REE-bearing primary minerals such as calcite, dolomite, apatite or sulphides. The solutions become increasingly enriched in Ba,

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F, SO, Sr, REE and Th, and precipitate REE phosphates if phosphate is available, or carbonates if phosphate is not sufficiently abundant. The REE mineralization in these environments tends to mirror that of the original carbonatite, i.e., fractionated in favour of the LREE.

The Emanya intrusion has abundant iron oxide veining and locally fluorite, indicating that some hydrothermal alteration has occurred and its REE concentrations are significantly enriched over those in the Main intrusion, although it is still LREE-dominated. Like the Emanya intrusion, many of the carbonatite dykes are significantly enriched in REE and fractionated in favour of the LREE.

At Lofdal, in addition to the LREE-enriched carbonatite-hosted mineralization, there is a late stage, structurally-controlled hydrothermal alteration that has resulted in HREE-rich mineralization in dynamic basement structures (see Figure 27 and Figure 29). These structural zones apparently acted as fluid pathways during mineralization and late-stage alteration in these structures introduced a HREE-rich mineral assemblage dominated by xenotime, and accompanied by zircon, rutile, apatite, fluorite and thorite. This is economically significant because it is the HREE that are the most valuable of the REE.

The current working hypothesis is that this HREE-rich hydrothermal alteration resulted from some combination of extended fractional crystallisation of the carbonatite magma and/or differential transport of the REE in exsolved hydrothermal fluids. Crystallisation of LREE-rich minerals early in the fractionation history could have resulted in a HREE-rich residual fluid phase, which escaped into selected structures during the later stages of crystallisation, resulting in the HREE-rich mineralization (Swinden and Burton, 2012). Depending on the chemistry of the hydrothermal fluids, fractionation of the HREE from LREE may also have occurred as a result of hydrothermal activity.

To date, the highly anomalous HREE enrichment at Lofdal has only been observed in selected structures associated with a complex series of alteration lithologies. The origin of the HREE enrichment is still uncertain. If it represents a late-stage fluid that evolved from extended fractionation of carbonatite magma, then the possibility exists that a plug of similar enrichment may be present in the subsurface. Alternatively, this HREE event may record expulsion of late hydrothermal fluids that re-distributed HREE already present in the rocks, in which case the primary targets will continue to be in the structures.

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9. Exploration

NMI began exploration on EPL 3400 in 2006. Exploration to date includes:
Regional and detailed exploration between 2006 and 2008 for copper and gold, targeting an IOCG model.
Regional assessment of the REE potential of the Lofdal Carbonatite Complex beginning in 2008 and continuing to the present.

9.1. Copper – Gold Exploration: 2006 – 2008

NMI was originally attracted to the Lofdal area as an IOCG target (Lobo-Guerrero, 2005). The model was predicated on the presence of copper sulphides associated with magnetite in the matrix of hydrothermal breccias and sulfidation accompanying magnetite and hematite in quartz veins possibly related to REE, Th, U and P bearing carbonatite dykes.

The IOCG exploration program consisted of:

A comprehensive structural and satellite mapping exercise over an area of more than $10,000\mathrm{km}^2$. The work was carried out under contract by the NPA Group of consultants, London, UK. Additional Landsat images were obtained and the Landsat thematic mapper™ images with two spectral bands (bands 5 and 7) in the Short-Wave Infrared (SWIR) were used to identify targets/outcrops of hydrothermal alteration and to delineate major structural features.
Targets generated by the NPA study, as well as other areas of interest, were systematically sampled by NMI personnel who collected a total of 2,371 rock grab samples. During the latter part of this phase of exploration in 2008, the first 255 samples were collected from the Lofdal Carbonatite Complex.
Areas returning anomalous Au values were sampled further, and a program of detailed geological mapping and trenching was undertaken to assess the most prospective targets. Three of these targets were tested by reverse circulation drilling in 2008 but no significant Cu-Au mineralization was found.
An orientation stream sediment geochemical program was undertaken by NMI personnel in the area of the former Lofdal Copper Mine but did not generate any significant targets.

Virtually all of this exploration was outside of the current area of interest for REE mineralization and few of the data are directly applicable or relevant to the current project. Where the 2006 to 2008 data are relevant to the REE exploration, they are noted in the following sections.

9.2. Regional Assessment of Rare Earth Element Potential

9.2.1. Geological and Lithogeochemical Survey

Lobo-Guerrero (2005) recognized and recommended to NMI the potential for REE mineralization associated with carbonatite dykes in the area. This potential was reinforced by the results of investigations by the GSN on the carbonatite. The company initiated an exploration campaign to test this potential towards the middle of 2008.

The initial regional field surveys of the Lofdal Carbonatite Complex were carried out by NMI personnel in two field campaigns from 2008 to February 2010. The aim was to systematically map and sample REE mineralization within the Lofdal Carbonatite Complex.

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The area of interest was defined by an airborne radiometric high (dominantly thorium) that NMI personnel interpreted to potentially define the extent of the intrusive complex. Sampling during these two campaigns covered roughly the northwestern half of the thorium anomaly, including the areas of known REE mineralization (Figure 44).

Note: Samples taken before early 2010 are indicated by white dots, samples taken after late 2010 are indicated by black dots; Background map shows Th radiometric counts from the 2010 airborne survey; Red and purple colours represent high values and define the extent of the Th anomaly associated with the Lofdal Carbonatite Complex. Intrusion (Carbonatite) shown in grey.
Source: Base map from New Resolution Geophysics, 2010
Figure 44 Distribution of regional lithogeochemical samples in the Lofdal area

In addition to the regional traversing, detailed sampling on approximately 50 m intervals was carried out over the Emanya intrusion and portions of the Main intrusion with 217 samples collected from the Emanya intrusion and 171 from the Main intrusion. A total of 3,680 grab samples were taken during these campaigns and the results were discussed in detail by Swinden and Siegfried (2011).

Surface grab samples continued to be a principal exploration tool on the property between 2011 and 2013. Regional geological mapping, coupled with regional airborne geophysics, continued to identify new alteration zones and carbonatite dykes which were systematically sampled to test their mineralizing potential and define drill targets. Systematic surface grab sampling was carried out to better define the mineralized systems throughout the area of interest (Figure 44). This sampling totalled approximately 1,900 additional samples. All surface grab samples were analysed at Actlabs in Ancaster, Ontario. The grab samples were analysed for the same suite of trace elements and were subjected to the same QAQC protocols as the drillhole core samples from Area 2B (detailed in Swinden and Siegfried, 2011) and Area 4 (detailed by Siegfried and Hall, 2012).

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The lithogeochemical surveys outlined the distribution of the REE mineralization at a district scale. They showed that there is considerable variation in the mineralization of the alteration systems – some being well mineralized and others containing no REE. Within the REE-mineralized zones, there is considerable variation in the grade of REE mineralization at both large and small scales. Mineralization is dominantly hosted by basement structures which have been altered and mineralized by hydrothermal fluids and intruded by carbonatite dykes. The linear spatial distribution of anomalous grab samples reflects the favourable structures, provides further evidence that only certain structures on the property contain significant HREE mineralization (Figure 45) and allowed these structures to be traced for considerable distances along strike. The regional geochemical surveys showed that the most favourable structures are notably enriched in the HREE and integration of surface lithogeochemistry with regional geophysical and geological studies resulted in the definition of nine priority exploration areas on the property (Figure 45 and Figure 46) were a primary tool in identifying drilling targets within these areas. Priority target areas have all been tested by drilling (see Item 10).

Source: Base map sourced from Landsat Geocover mosaic, 2011
Note: Main intrusion shown in grey for reference.
Figure 45 Priority exploration areas defined by Dy in surface lithogeochemistry samples

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Source: Base map sourced from Landsat Geocover mosaic, 2011
Note: Main intrusion shown in grey for reference.
Figure 46 Priority exploration areas defined by HREE/TREE ratio in surface lithogeochemistry samples

9.2.2. Remote Sensing and Regional Geophysics

The results of the regional remote sensing and geophysical surveys were previously reported by Swinden and Siegfried (2011) and are summarized in Table 12. NMI has made extensive use of remote sensing data in interpreting the geological relationships on the property and identifying priority exploration targets.

Table 12 Summary of remote sensing and regional geophysical surveys and interpretations

Method	Contractor	Objectives	Results
HyMap data analysis – 126 bands from 0.4 to 2.5μm t 4.6m resolution	NamibGeoVista (2010)	Identify hyperspectral signatures characteristic of carbonatite dykes and plugs.	Successfully imaged larger carbonatite and phonolite dykes.
ASTER	NPA Fugro (2010)	Interpret geology, delineate carbonatite bodies.	Data reflect mineralogical characteristics of some basement lithologies. Not effective in targeting individual dykes or intrusions.
Structural interpretation using Landsat/HyMap	NPA Fugro (2010)	Identify major structural features.	Identified and mapped major first order shear zones and other second order structures, some of which are mineralized; provides a structural context for the basement and the Lofdal Carbonatite Complex intrusions.
Integration of remote sensing data with	NPA Fugro (2010)	Identify priority target areas.	First pass at regional target definition Primary targets

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Method	Contractor	Objectives	Results
geophysics and geochemistry			identified using magnetic, calcite/iron spectral signatures, structural control, Th radiometric signature and HREE in proximity. Secondary targets as above but lacking strong magnetic signature.
High resolution airborne radiometric and magnetometer survey; helicopter-borne, 75 m line spacing, 3760 line km at 315°Az	NPA Fugro (2010)	Achieve better resolution of radiometric and magnetic features.	Confirmed the radiometric signatures of the carbonatite intrusions and dykes. Confirmed that individual dykes can be traced. Confirmed interpreted structural trends.
Regional ground radiometric, magnetic, gravity profiles (5 lines)	Greg Symons Geophysics (2010)	Test geophysical signatures of intrusions and airborne geophysical targets, test whether gravity identifies known intrusions and can identify buried bodies.	Radiometric and magnetic data respond to individual dykes, consistent with airborne results. Gravity was inconclusive regarding response of carbonatite intrusions or presence of additional carbonatite bodies at depth.

The 2010 high-resolution airborne survey provides high-resolution information that correlates well with existing geological, geophysical and lithogeochemical data for the area. In particular:

It confirmed the contrasting radiometric signatures of the Main and Emanya intrusions.
The resolution was sufficiently high to confirm and enhance the mapping of REE-bearing structures.
It confirmed interpreted structural trends and allowed more detailed mapping of major structures across the area, particularly where these structures appear to have served as conduits for carbonatite intrusion and mineralization.
In summary, the regional exploration identified multiple, high quality REE target areas and demonstrated significant potential for discovery of deposits of REE associated with the Lofdal Carbonatite Complex. Key results of the regional exploration include:
Recognition of the district scale thorium anomaly which provides a first order regional target for REE exploration in the Lofdal area.
Dramatic expansion of the number and extent of known carbonatite dykes and related alteration zones and documentation of their geological characteristics and associated REE.
Recognition that carbonatites and the associated rocks are extensively hydrothermally altered and variably mineralised with REE.
Recognition that the HREE-rich mineralization is structurally controlled and that certain structures are preferentially enriched in HREE mineralization.
Geological and geophysical characterisation of two intrusive plugs in the centre of the complex.

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High resolution geophysical characterisation of the area, interpretation of the regional structural setting of the complex, and recognition of hyperspectral and geophysical signatures that characterize carbonatite dykes and plugs.
Identification of a number of high-priority target areas for detailed exploration with new targets being generated as field work and compilations continue.

9.2.3. Regional Geological Mapping

Published geological maps for the Lofdal area are at a scale of 1:250,000 (GSN, 2006). This is too broad a scale to be useful for property-scale investigations. Accordingly, detailed mapping of the core of the Lofdal Carbonatite Complex was initiated in 2010 and continued through the latter part of 2013. The mapping was carried out by geologists of South Africa-based Remote Exploration Services Ltd. (RES) on 100 m spaced traverse lines. Extensive use was made of hyperspectral data, which was recognized to closely reflect basement lithologies and areas of hydrothermal alteration. This allowed accurate extrapolation of rock units and alteration zones between and beyond traverse lines. The mapping began with detailed mapping in Areas 2 and 4 to support the planned trenching and drilling in 2010 and then expanded to include the area of the intrusive core of the Lofdal Carbonatite Complex. The mapping continued in 2012 and 2013 in outlying areas of the property and by the end of 2013, the entire extent of the thorium anomaly that defines the exploration area of interest had been geologically mapped. The geological map of the property is shown in Figure 20.

Detailed mapping has contributed to the understanding of the geology and therefore the exploration at Lofdal in several important ways:

Clarified the nature and distribution of basement lithologies.
Clarified the distribution of intrusive lithologies related to the Lofdal Carbonatite Complex. It demonstrated that the Main intrusion is dominantly syenite at the current level of exposure, but, close to the contact with the underlying white calcite carbonatite, it identified and mapped the distribution of two additional nepheline syenite plugs to the southwest of the Main intrusion and mapped the extent of the Emanya carbonatite intrusion.
Mapped the distribution and intensity of phonolite and carbonatite dykes related to the complex.
Mapped the distribution of Lofdal breccias, showing that they are widespread along the axis of the Lofdal intrusions. The intrusive axis of the Lofdal Carbonatite Complex, as defined by nepheline syenite and carbonatite intrusions and related breccia, occupies a strike length of more than 6 km.
Showed that most intrusions have a halo of fenitisation and that hydrothermal alteration in the form of albitisation can be mapped out along some major structures for several kilometres along strike. The combination of alteration, anomalous HREE geochemistry and radiometric anomalies related to Th have been important in identifying priority exploration targets in the complex.
Traced hydrothermal alteration zones along strike well beyond their previously known extent and identified a number of new albitite-carbonate alteration systems, particularly in the northeast and southeastern parts of the property, that comprise exploration targets for further concentrations of HREE.

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9.3. Target Exploration in Area 2B

The regional assessment of the REE potential of the Lofdal Complex led to an initial selection of Areas 2 and 4 for more detailed exploration (Figure 45). The areas were chosen on the basis of the presence of carbonatite dykes and albitic/carbonatitic alteration zones with significant widths in outcrop, a high relative proportion of samples with anomalous REE values (particularly high HREE/TREE ratios), and geophysical / hyperspectral signatures potentially indicating the presence of more extensive zones of carbonatite. The initial focus of this work was the prominent carbonatitic and albitic alteration zone in Area 2B. The exploration of this zone comprised detailed geological mapping, lithogeochemistry and trenching and was largely completed during 2010. This work was described in detail by Swinden and Seigfried (2011) and is summarized in the following sections:

9.3.1. Geological Mapping and Lithogeochemistry

Area 2B contains a segment of a regionally significant, ±3 km long, carbonatite dyke – alteration system in which almost 50% of surface samples contain more than 0.5% TREE and more than 25% of samples contain more than 0.1% HREE. Regional mapping defined the overall setting of the zone and detailed mapping in Area 2B defined a mineralized alteration zone that is exposed over significant widths (up to 70 m at surface) along a strike length of more than 650 m. The detailed mapping showed that the Area 2B alteration/carbonatite zone strikes approximately 060° and dips steeply to the SE, cutting the structural grain of the basement at a high angle. The basement comprises mainly interlayered amphibolite and quartzo-feldspathic gneiss, locally cut by coarse grained pegmatite. Abundant carbonatitic dykes in this area vary widely in orientation but there is a dominant northeasterly strike similar to the mineralized alteration zone.

Detailed mapping showed that the alteration is dominantly expressed as an early pervasive albitization, which has been brecciated and overprinted by an albite-carbonatite-mica assemblage. There are abundant breccias in which angular albitite fragments are surrounded by carbonatite, indicating that the alteration zone was a brittle fracture zone throughout its active history.

The zone was initially targeted by the results of regional grab sampling. A second round of more detailed surface sampling resulted in a surface suite of approximately 50 grab samples. These samples were intended to be broadly representative of the mineralized zone but were not taken on a systematic grid pattern so do not provide a complete picture of the grade distribution in the surface exposure of the alteration zone. The second round of sampling confirmed the anomalous nature of the alteration system overall and the fact that the REE are very unevenly distributed within the alteration. The five highest-grade analyses from these samples are presented in Table 13.

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Table 13 Analyses of the five highest-grade surface samples in Area 2B

Sample	La ppm	Ce ppm	Pr ppm	Nd ppm	Sm ppm	Eu ppm	Gd ppm	Tb ppm	Dy ppm	Ho ppm	Er ppm	Tm ppm	Yb ppm	Lu ppm	Y ppm	HREE %	TREE %
NLOFR1636	1 630	2 560	243	1 060	765	320	954	182	994	217	544	82.3	417	61.5	4 840	0.86	1.48
NLOFR1653	1 035	1 640	149	526	301	174	648	163	945	201	469	67.1	319	43.9	4 840	0.78	1.15
NLOFR1652	1 640	2 430	240	754	237	142	542	138	874	166	400	44.2	205	24.6	4 220	0.67	1.20
ERN15349	1 470	2 060	186	574	254	120	392	81	529	110	321	44.4	265	35.5	3 070	0.49	0.95
NLOFR1618	862	1385	163	805	764	246	730	99	456	75	211	30.0	198	27.8	2 390	0.44	0.84

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9.3.1.1. Trenching

25 trenches (Figure 47) were dug across the Area 2B alteration zone to determine the distribution and geological setting of the REE in two dimensions. Trenching was carried out using a Bell 315J backhoe/loader (Figure 48). Bedrock is typically within 0.5 m of surface, although in parts of some trenches, up to 1.5 m of soil or calcrete cover was encountered. The trenches were then cleaned by hand in preparation for mapping and sampling (Figure 48) and metre waypoints were measured and marked by aluminium tags.

Source: Swinden, 2020
Figure 47 Trenches on the Area 2B Zone (heavy brown lines) – trenches listed in Table 14 are labelled

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Table 14 Locational information for trenches on the 2B Zone – WGS84; UTM Zone 33S.

Trench ID	Easting (m)	Northing (m)	Elevation (m)	Azimuth (°)	Length (m)
NLOFTR2B001	467 013.9	7 754 568.9	957.5	135	77
NLOFTR2B002	467 032.8	7 754 582.8	957.3	135	76
NLOFTR2B003	467 046.5	7 754 596.0	956.6	135	79
NLOFTR2B004	467 063.3	7 754 602.5	956.4	135	87
NLOFTR2B005	467 077.7	7 754 619.1	955.3	135	53
NLOFTR2B006	467 102.3	7 754 628.1	955.1	135	53
NLOFTR2B007	467 111.1	7 754 648.6	953.7	135	60
NLOFTR2B008	467 121.7	7 754 667.5	952.4	135	74
NLOFTR2B009	467 126.0	7 754 694.3	949.9	135	100
NLOFTR2B010	467 154.2	7 754 698.8	949.4	135	81
NLOFTR2B011	467 174.4	7 754 710.6	947.9	135	78
NLOFTR2B012	467 189.9	7 754 725.0	947.2	135	92
NLOFTR2B013	467 235.2	7 754 788.1	945.1	135	73
NLOFTR2B014	467 300.3	7 754 827.7	949.5	135	31
NLOFTR2B015	467 307.0	7 754 852.8	950.4	135	55
NLOFTR2B016	467 317.9	7 754 865.5	951.4	135	58
NLOFTR2B017	467 332.3	7 754 885.3	951.6	135	104
NLOFTR2B018	467 346.3	7 754 895.3	952.9	135	145
NLOFTR2B019	467 364.2	7 754 910.1	953.3	135	122
NLOFTR2B020	467 382.9	7 754 923.9	953.7	135	105
NLOFTR2B021	467 403.6	7 754 932.7	953.1	135	75
NLOFTR2B022	467 415.3	7 754 953.9	951.1	135	72
NLOFTR2B023	467 431.9	7 754 966.7	949.4	135	60
NLOFTR2B024	466 982.1	7 754 542.0	956.7	135	53
NLOFTR2B025	466 996.2	7 754 561.2	957.0	135	60

Source: Swinden, 2010

Figure 48 A – digging trenches on Area 2B with a backhoe, B – cleaning trenches in preparation for sampling and mapping

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All trenches were geologically mapped prior to sampling and total average count radiometric readings were taken for each sample interval using a RadEye personal radiation detector (PRD). Total count K, U, and Th concentrations with a hand-held spectrometer and magnetic susceptibility readings were taken at 1 m intervals.

The trenches were sampled using a hand-held diamond saw to make two parallel cuts approximately four cm apart. Calcrete coatings were removed by hammer prior to sampling. The trenches were then continuously sampled with hammer and chisel by removing the rock between the saw cuts.

The results of trench mapping confirm the fine structure of the carbonatite/alteration system. The mineralized zone is characterised by carbonatite dykes of variable width associated with variably altered schists and gneisses. The alteration is characteristic of the mineralized zones and consists of variable, locally intense, albitization, carbonatization and local concentrations of phlogopite. Channel samples were taken throughout the entire length of the trenches to test whether mineralization is restricted to carbonatites or is related to hydrothermal alteration and occurs in other lithologies as well. The favourable zones of carbonatization, alteration, and mineralization are well outlined by both carbonatite intensity and by total count radiometric signature.

The results of channel sampling and analysis of trench samples were presented in detail by Swinden and Seigfried (2011) and the best intersections are summarized in Table 15. The alteration zone is generally characterized by significantly elevated REE contents as well as MnO and P₂O₅, Th, U, Ba, Nb and Zr. Similar patterns of enrichment are seen in the alteration zone in other trenches.

The alteration system and its associated radioactive and geochemical anomalies was encountered to varying degrees in all trenches in Area 2B. However, the assays show that the zone is not consistently mineralized along its length. The best mineralized intersections in terms of both grade and width were encountered in Trenches NLOFTR2B001, 005 and 006. The assay results for these mineralized zones are illustrated in Table 15. Quoted intervals are horizontal, not true widths. In all three trenches, REE mineralization is FREE-enriched and developed over widths of several metres.

This trench sample information contributed to the mineral resource model by providing a position to which the mineralised zone can be extended to surface from the shallowest drillholes. The grades of the samples were not used in the block model grade estimation, however, due to biases considered to be the result of very near surface enrichment with very limited vertical extent.

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Table 15 Assays from best trench intersections -NLOFTR001, NLOFTR005 and NLOFTR006, Area 2B

Trench	From	To	Width (m)	La ppm	Ce ppm	Pr ppm	Nd ppm	Sm ppm	Eu ppm	Gd ppm	Tb ppm	Dy ppm	Ho ppm	Er ppm	Tm ppm	Yb ppm	Lu ppm	Y ppm	HREE %	TREE %
Trench 1	26	28	2	2 338	3 455	363	1 758	1 287	472	1574	238	1 134	182	382	44	244	32	4 479	0.88	1.80
	39	42	2	22	40	5	29	45	26	103	23	155	34	98	14	85	12	1 177	0.17	0.19
Trench 5	33	44	11	817	1 230	117	483	268	130	502	94	545	97	245	32	184	25	2 676	0.45	0.74
	36	40	4	1 129	1 669	159	675	451	233	919	167	937	159	384	47	260	35	4 295	0.74	1.15
	11	12	1	281	459	45	172	51	25	86	17	107	19	47	6	36	5	511	0.09	0.19
Trench 6	29	45	16	648	961	93	407	236	96	355	59	329	57	144	19	113	16	1 595	0.28	0.51
	32	36	4	561	887	92	419	313	132	495	81	434	72	175	23	133	19	1 947	0.35	0.58
	39	45	6	1 236	1 802	170	736	382	152	561	96	536	92	234	30	164	22	2 662	0.45	0.89

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9.4. Target Exploration in Area 4

9.4.1. Geological Mapping and Surface Sampling

Geological mapping in Area 4 identified a prominent zone of carbonatization and REE mineralization trending approximately east to west across the area which is particularly enriched in HREE (Figure 30 and Figure 31). The alteration zone is readily mapped as a zone in which both carbonatites and alteration lithologies are coloured dark brown by the weathering of iron oxides. The zone has both significant width and strike length and is associated with a prominent Th radiometric anomaly. Detailed sampling reported by Swinden and Siegfried (2011) showed that a large proportion of samples from the core of the alteration zone returned very strong HREE-enrichment and also identified an area in the southwest quadrant of the area which returned a large number of LREE-enriched samples. These samples were intended to be broadly representative of the mineralized zone but were not taken on a systematic grid pattern so do not provide a complete picture of the grade distribution in the surface exposure of the alteration zone. Assays from selected grab samples confirmed that the mineralization is unevenly distributed in the alteration zone. Samples from these two areas returned up to 4.69% TREO with 96.3% HREE enrichment (%HREE/%TREE) and 5.82% TREO with 2.8% HREE-enrichment, respectively.

9.4.2. Ground Geophysics

Ground radiometric and gradient induced polarization/resistivity surveys were carried out across the Area 4 alteration zone on 25 m spaced lines. The radiometric survey was carried out by Greg Symons Geophysics in 2010. The gradient survey was carried out by Remote Exploration Services in 2011 and extended by Greg Symons Geophysics in 2012. The results of this work were described in detail by Siegfried and Hall (2012) and are summarized in Table 16 and illustrated in Figure 49.

Note: North is upwards, grid line spacing is 400 m for scale
Source: RES, 2011; Symons Geophysics, 2012
Figure 49 Area 4 grid showing ground geophysical coverage

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Table 16 Summary of geophysical surveys in Area 4
| Radiometric Survey | Continuous walking mode = station spacing ~2m | Alteration zone follows a prominent ENE-striking Th anomaly, follows 1storder sinistral shear |
| --- | --- | --- |
| Gradient array induced polarization/resistivity surveys | RES- 25 m line spacing, 25m station spacing and a-spacing; GSG – 50 m linespacing, select lines of pole-dipole array with a=25m | Good agreement between resistivity and inverted 3D PDP. Central low resistivity belts coincide with alteration zone |

Geophysical surveys along the Area 4 trend identified a number of priority targets for further investigation both on the central trend that is interpreted to represent the NE extension of the Area 4 fault zone, as well as off the corridor trend. The on-axis targets reflect combinations of low resistivity (suggesting fault structures, particularly cross-cutting or splay structures), strong alteration (mapped) and moderate to strong HREE enrichment in surface samples, and correlation with high Th in ground and airborne geophysics.

9.4.3. Trenching

Twenty trenches were dug across the Main zone in Area 4 to locate and sample the mineralised zone precisely at surface and examine its alteration and geological relationships to the surrounding rocks (Figure 50 and Table 17).

This information contributed to the mineral resource model by providing a position to which the mineralised zone can be extended to surface from the shallowest drillholes. The grades of the samples were not used in the block model grade estimation, however, due to biases considered to be the result of very near surface enrichment with very limited vertical extent.

Note: Trenches are shown with brown lines. The Area 4 alteration zone is indicated by dark grey (intense alteration) and light grey (moderate alteration). Samples are keyed to the ratio HREE/TREE. Geological and sample legends same as Figure 47.
Source: Swinden and RES, 2012
Figure 50 Location of Trenches in Area 4

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The trenches were dug using a JCB backhoe and the width of the trenches was determined by the excavator bucket (Figure 51), which was approximately 1 m. Trench endpoints and midpoints were located by GPS. Bedrock is typically within 50 cm from surface although in parts of some trenches up to 1.5 m of soil or calcrete cover was encountered. The trenches were cleaned by hand in preparation for mapping and sampling and metre waypoints were measured and marked by aluminium tags.

All trenches were geologically mapped prior to sampling and total average count radiometric readings were taken for each sample interval using a RadEye personal radiation detector (PRD). Total count K, U and Th concentration readings were taken at 1 m intervals using a hand-held spectrometer and magnetic susceptibility.

Source: Hall, 2012

Figure 51 Examples of trenches in Area 4

Table 17 Locational information for trenches in Area 4. WGS84, UTM Zone 33S

Trench ID	East_Start (m)	North_Start (m)	East_End (m)	North_End (m)	Length (m)
NLOFTR4001	470 136	7 753 404	470 111	7 753 489	88
NLOFTR4002	470 183	7 753 421	470 164	7 753 494	99
NLOFTR4003	470 204	7 753 445	470 192	7 753 506	75
NLOFTR4004	470 245	7 753 415	470 223	7 753 516	62
NLOFTR4005	470 297	7 753 474	470 280	7 753 539	104
NLOFTR4006	470 342	7 753 486	470 319	7 753 578	67
NLOFTR4007	470 395	7 753 497	470 371	7 753 589	95
NLOFTR4008a	470 442	7 753 510	470 425	7 753 577	95
NLOFTR4008b	470 422	7 753 585	470 421	7 753 592	70
NLOFTR4009a	470 496	7 753 513	470 476	7 753 576	7
NLOFTR4009b	470 473	7 753 585	470 465	7 753 614	66

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Trench ID	East_Start (m)	North_Start (m)	East_End (m)	North_End (m)	Length (m)
NLOFTR4010	470 520	7 753 512	470 494	7 753 608	30
NLOFTR4011	470 578	7 753 480	470 548	7 753 612	135
NLOFTR4012	470 618	7 753 532	470 589	7 753 642	113
NLOFTR4013	470 651	7 753 593	470 633	7 753 664	74
NLOFTR4014	470 084	7 753 422	470 072	7 753 502	80
NLOFTR4015	470 040	7 753 370	470 020	7 753 500	131
NLOFTR4016	469 996	7 753 391	469 979	7 753 524	134
NLOFTR4017a	469 943	7 753 344	469 935	7 753 398	54
NLOFTR4017b	469 934	7 753 412	469 918	7 753 504	94
NLOFTR4018	469 877	7 753 408	469 869	7 753 483	75
NLOFTR4019	469 821	7 753 420	469 812	7 753 507	86
NLOFTR4020a	469 785	7 753 373	469 783	7 753 393	21
NLOFTR4020b	469 781	7 753 402	469 781	7 753 415	13
NLOFTR4020c	469 778	7 753 426	469 769	7 753 495	72

All trenches crossed the Main zone of REE mineralization and alteration, and this zone is marked in each trench by anomalous radioactivity, visible evidence of alteration and geochemical anomalies in the HREE and Y, $\mathrm{HREE + Y / TREE + Y}$, and $\mathrm{P_2O_5}$. In almost all cases, the geochemical and radiometric anomalies coincide closely with the mapped extent of the alteration and mineralization. Representative assays are given in Table 18. The detailed sampling from the trenches confirms the preliminary observations of significant grades (\%TREE+Y) accompanied by very high levels of HREE-enrichment (66% to $>90\%$) over a continuous strike length of up to $650~\mathrm{m}$.

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Table 18 Representative analyses from trench samples, Area 4

Trench	From	To	Width (m)	La ppm	Ce ppm	Pr ppm	Nd ppm	Sm ppm	Eu ppm	Gd ppm	Tb ppm	Dy ppm	Ho ppm	Er ppm	Tm ppm	Yb ppm	Lu ppm	Y ppm	HREE %	TREE %
NLOFTR4001	52	54	2	67	159	22	110	67	30	110	22	141	28	77	11	64	9	939	0.14	0.19
NLOFTR4004	78	94	16	46	82	11	48	67	41	196	46	308	67	190	28	162	23	1 993	0.31	0.33
	88	90	2	76	155	19	91	161	107	520	122	866	187	533	77	447	64	5573	0.85	0.90
NLOFTR4005	45	57	12	26	47	8	41	78	52	300	80	568	128	376	53	298	43	4459	0.64	0.66
	49	52	3	20	37	9	60	152	110	683	190	1 379	317	938	132	738	106	11 084	1.57	1.60
NLOFTR4006	39	66	27	132	232	26	101	52	31	170	47	325	76	220	33	191	28	2483	0.36	0.41
	48	49	1	295	518	58	230	183	129	714	199	1 360	304	843	123	701	100	8563	1.30	1.43
	58	59	1	42	93	14	86	169	121	729	211	1 520	353	1 030	153	884	127	12 230	1.74	1.78
NLOFTR4011	112	130	18	154	298	33	127	74	46	264	73	547	127	379	61	366	55	3765	0.57	0.64
	113	114	1	162	305	33	132	110	74	465	125	979	227	659	103	600	87	7342	1.07	1.14
	118	119	1	43	87	11	46	67	58	360	104	828	194	570	90	541	79	5594	0.84	0.87
	126	129	3	301	565	63	244	180	127	801	238	1 783	420	1 295	213	1 279	193	12 483	1.88	2.02
NLOFTR4013	35	43	8	304	508	51	176	53	24	119	30	209	47	140	22	139	21	1314	0.21	0.32
	37	38	1	180	318	36	150	99	61	376	105	765	176	513	76	468	68	5137	0.77	0.85

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10. Drilling

Several campaigns of diamond drillhole drilling were undertaken from 2010 to 2020, including detailed mineral resource drilling in Areas 2B and 4, and exploration drilling on a wide variety of targets throughout the ML. Drilling procedures and results for the earlier campaigns have previously been reported in detail (Swinden and Siegfried, 2011; Siegfried and Hall, 2012; Dodd et al., 2014).

The drilling is discussed in several sections:

2010 and 2011 – exploration drilling in Area 2B, which eventually contributed to the mineral resource estimate in this area;
2011 and 2012 – mineral resource drilling in Area 4, which led to the definition of the first mineral resource on the property;
2020 – mineral resource drilling in Areas 2B and 4 which defined and expanded the existing resource in Area 4 and led to the definition of the Area 2B mineral resource;
2010 to 2020 – exploration drilling which has tested multiple targets but has not led to definition of further mineral resources; and
2023 – a reverse circulation (RC) campaign undertaken as an in-fill programme and to prove extension of the mineralisation along strike and at depth.

10.1. Area 2B, 2010 and 2011 Drilling

The 2010 drilling (Holes NLOFDDH2B001 – 013) was previously reported in detail by Swinden and Seigfried (2011) and is summarized below. The initial drilling was carried out in October and November 2010 and managed by GeoAfrica Prospecting Services. Drilling procedures were governed by a Standard Operating Procedures (SOP) Manual developed for the project by GeoAfrica Prospecting Services and approved by Namibia Rare Earths Ltd. Diamond core drilling on the Area 2B target in 2010 comprised an exploratory phase of 13 drillholes totalling 154.5 m.

Table 19 presents a summary of information and procedures with respect to this drilling campaign. Figure 52 shows the location of the drillhole collars.

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Table 19 Summary of drilling procedures for the 2010 drilling campaign in Area 2B

Purpose of Drilling	Investigate the lateral and vertical extent, geology and grade of the Area 2B alteration zone.
Drill Equipment	Two Atlas Copco Christianson CS14 diamond drill rigs.
Core diameter	NQ (47.6 mm).
Hole Characteristics	All holes drilled at 323° to 330° azimuth and at 50° to 60° degrees dip. See Table 10.1.2 for locational and directional information. The alteration zone was found to dip at ~50° to 60° and intersections were at depths of 25 m to 50 m.
Rig Set-up	Checked by geologist prior to drilling. Holes collared with HQ to about 6 m depth.
Casing	All holes cased to bedrock and casing left in holes.
Site rehabilitation	Area of holes was rehabilitated, and a concrete plinth constructed showing UTM coordinates, drillhole number, hole depth, month and year. Top of casing sealed with a closed galvanized iron tube riveted to casing.
Drillhole orientation	All holes were downhole surveyed with a Reflex EZ-TRC system to determine drillhole azimuth and inclination.
Collar Locations	Collars of completed holes were DGPS surveyed by a professional surveyor (WGS84/UTM 33S).
RQD	Run by run recoveries monitored and recorded on a standard rock quality designation (RQD) sheet. RQD and recoveries were captured in the field.
Core Marking	Metre marking of core was done on site. Depth corrections were done by identifying drilling breaks and rejoining/remeasuring the core. Zones of core loss were recorded. High points of contacts/layering were used for orienting and marking the rejoined core. An orientation line was marked on the entire length of the core with arrows added frequently to indicate downhole direction
Core Handling	Core was secured in boxes and transported securely by vehicle from the drill site to the core yard in Khorixas.
Core Logging	All holes were logged and sampled by the same qualified geologist. Logging was carried out on standard company logging forms and imported into spreadsheets and a Microsoft Access database. Several holes were radiometrically logged using a RadEye PRD gamma scintillometer. All holes were photographed when wet and then marked with indelible pens for sampling.
Downhole logging contracted to Terratec Geoservices (Namibia)	Three arm caliper/gamma as an initial test for the drillhole integrity.
	Dual compensated density/gamma side walled tool which provided a measure of the density of the material surrounding the drillhole
	Magnetic susceptibility.
	Acoustic televiewer provided a 360° acoustic image of the drillhole with directional information (dip and azimuth) of fractures and layers.
	Down hole spectrometric record of the U, Th and K concentration of the rock.

Four additional diamond drillholes were drilled in Area 2B in August 2011 (Holes NLOFDDH2B014 - 017). These holes were drilled as part of the 2011 drilling campaign that included the resource drilling in Area 4. Drilling procedures were identical to the Area 4 drilling, governed by the same SOPs, and QAQC for these holes was integrated in and accounted for by the Area 4 QAQC described by Siegfried and Hall (2012). The intention of these holes was to test a part of Area 2B that had returned the best values in the previous year. The results were not deemed to be sufficiently encouraging to continue at that time.

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Additional infill sampling was undertaken in 2020 on intervals from five drillholes completed during 2010 and 2011. These samples were assayed as part of the 2020 campaign, and sampling, sampling preparation and assaying were carried out in accordance with the SOP in effect in 2020. QAQC was accounted for as part of the 2020 drilling campaign. Table 20 details the intervals sampled.

Table 20 Infill samples from 2010-2011 drillholes sampled and assayed during 2020

Hole ID	From (m)	To (m)	Number of samples
NLOFDH2B007	50.00	55.00	5
NLOFDH2B008	60.30	66.00	5
NLOFDH2B012	17.35	22.70	6
	57.00	64.00	7
NLOFDH2B013	13.00	22.00	9
	67.00	73.00	6
NLOFDH2B015	66.00	73.00	7
	145.00	149.20	5

Note: Geological legend as for Figure 47.
Source: Swinden, 2014
Figure 52 Location of drillhole collars in Area 2B

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10.2. Area 4 Mineral Resource Drilling, 2011 and 2012

Drilling on Area 4 started in June 2011 and continued until the end of April, 2012. The drilling was managed on-site by Remote Exploration Services (RES) of Cape Town, South Africa, under Standard Operating Procedures dated May 11, 2011 and developed for this project by RES. A total of 101 diamond drillholes were completed, of which 93 comprised a systematic grid-based assessment of alteration and mineralization in the Main zone and eight were drilled down dip on the Main zone with larger HQ diameter core to obtain material for metallurgical test work (Figure 53). Drilling in this campaign totalled 11,783.6 metres.

The 2011 and 2012 drilling were previously reported in detail by Siegfried and Hall (2012) and Dodd et al. (2014).

Table 21 Summary of drilling procedures for the 2011-2012 drilling campaign in Area 4

Purpose of Drilling	Investigate the lateral and vertical extent, geology and grade of the Area 4 alteration zone. Resource drilling and metallurgical sampling.
Drilling Contractor	JGM Drilling and Exploration Namibia.
Drill Equipment	Two CF90 platform-mounted diamond drill rigs.
Core diameter	NQ (47.6 mm). Drilling utilized standard and triple tube 4 9/16" HQ and NQ core barrels.
Hole Characteristics	All exploration and resource holes drilled at approximately 345° azimuth and at -55° dip, metallurgical holes at -40° dip. See Table 10.1.2 for locational and directional information. The alteration zone dips southerly at between 45° and 60°. The systematic grid drilling intersected mineralization to vertical depths of between 150 m and 200 m. Four deep holes intersected the zone at 250 m and 300 m vertical depth.
Rig Set-up	Location and inclination checked by geologist prior to drilling. Holes collared with HQ to about 6 m depth.
Core Recovery	>90%; Drillholes with inadequate recovery were re-drilled by the contractor with suffix “B” added to the original drillhole number.
Casing	All holes cased to bedrock and casing left in holes.
Site rehabilitation	Area of holes was rehabilitated, and a concrete plinth constructed showing UTM coordinates, drillhole number, hole depth, month and year. Top of casing sealed with a closed galvanized iron tube riveted to casing.
Drillhole orientation	All holes were downhole surveyed with a Reflex EZ-Shot system to determine drillhole azimuth and inclination.
Collar Locations	The collar positions and elevation were surveyed by a professional surveyor with a real-time kinematic (“RTK”) GPS. (WGS84/UTM 33S).
RQD	Run by run recoveries monitored and recorded on a standard rock quality designation (RQD) sheet. RQD and recoveries were captured in the field.
Core Marking	Metre marking of core was done on site. Depth corrections were done by identifying drilling breaks and rejoining/remeasuring the core. Zones of core loss were recorded. High points of contacts/layering were used for orientating and marking the rejoined core. An orientation line was marked on the entire length of the core with arrows added frequently to indicate downhole direction.
Core Handling	Core was transported securely from the drill site to the exploration camp on a daily basis.
Core Logging	Geological and geophysical logging was carried out by RES and NMI geologists and followed a comprehensive protocol. All holes were radiometrically logged using a RadEye PRD gamma scintillometer and

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	magnetically logged using a hand-held magnetic susceptibility meter. All holes were photographed when wet and then marked with indelible pens for sampling.
Downhole logging contracted to Terratec Geoservices (Namibia)	Three arm caliper/gamma as an initial test for the drillhole integrity.
	Dual compensated density/gamma side walled tool which provided a measure of the density of the material surrounding the drillhole.
	Magnetic susceptibility
	Down hole spectrometric record of the U, Th and K concentration of the rock.

Figure 53 Plan showing collars of the 101 drillholes drilled in 2011 and 2012 in Area 4

Note: Geological legend as for Figure 47
Source: Swinden, 2014

10.3. Areas 4 and 2B Mineral Resource Drilling, 2020

A campaign of drilling was undertaken on the Lofdal ML during 2020 with the objective of defining a mineral resource at Area 2B and expanding the existing mineral resource at Area 4. Gecko Exploration (Pty) Ltd was contracted to oversee and manage the drill program.

Drilling was undertaken in Area 4 between late February 2020 and early December 2020. The objective of this work was to extend the mineral resource in this area, particularly along strike to the west and to vertical depths of greater than 200 m. A total of 56 diamond drillholes were completed totalling 10,162.1 metres. Collar locations are shown in Figure 54.

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Note: Geological legend as for Figure 47. Red, black and blue squares same as Figure 53. Red triangles are 2020 drilling.
Source: Swinden, 2014
Figure 54 Plan showing drillhole collars in Area 4

Drilling was also undertaken in Area 2 from August to October 2020, with the objective of expanding on the previous drilling to attempt to define a mineral resource in this area. A total of 29 diamond drillholes were completed totalling 4,400.4 m. Collar locations are shown in Figure 52.

10.3.1. Area 4 and 2B Diamond Drilling Procedures

Diamond drilling during 2020 was governed by Standard Operating Procedures developed for Namibia Critical Metals Inc. by Gecko Exploration (Pty) Ltd. and approved by MSA. Diamond core drilling was undertaken by Günzel Drilling of Namibia with two diamond drill rigs; an Atlas Copco CS14 and an Atlas Copco CS1000. The first three to six metres of each hole were drilled with HQ diameter (63.5 mm) effectively collaring the hole and allowing casing to be inserted. The remainder of each hole was usually completed at NQ diameter core (47.6 mm). All holes in Area 2B were drilled towards an azimuth of 310° to 315° at dips between 60° and 68°. Those in Area 4 were drilled towards an azimuth of approximately 340° at dips between 55° and 71°.

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All of the 2020 diamond drillhole cores were orientated and an orientation line was marked on the core to guide structural measurements.

10.3.2. Core Recovery

The drill advance was marked by a Günzel Drilling technician on depth blocks after each drill run. Metre marking of the core as well as rock quality designation (RQD) and core recovery measurements were undertaken at the drill site by Gecko technicians. Orientation lines were drawn on the core with arrows indicating the down-dip direction (Figure 55). Core recovery was generally very good (>95 %). Core boxes were transported by vehicle daily from the drill site to the logging facility at the Lofdal base camp. Core was carefully loaded and ratchet strapped for transportation.

At the Lofdal camp, the core was logged radiometrically using a RadEye PRD gamma scintillometer and visibly altered sections were checked with an Olympus Delta 50 portable XRF to ensure that all mineralized sections were identified and sampled. Following geological logging, the core was sampled for assay.

Source: Ellmies, 2020
Figure 55 Drillers' metre marks, measured metre marks and orientation lines on uncut core

10.3.3. Collar and Downhole Surveys

The drillhole collar positions were pegged by a geologist using a handheld global positioning system (GPS) set within WGS84, UTM Zone 33S coordinate system. The senior geologist then verified the correct orientation and inclination of the rig derrick prior to drilling. After the completion of each hole, Günzel Drilling carried out downhole surveys using a Reflex EZ-Trac survey tool determining the downhole orientation, i.e., the dip and azimuth. The collar positions and elevation were surveyed by Greg Symonds Geophysics with a real-time kinematic (RTK) GPS. The drillhole collars were marked with a concrete beacon recording the relevant details of each hole on a metal plate (Figure 56).

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Source: Witley, 2021 (left), Witley, 2023 (right)

Figure 56 Concrete plinth over capped diamond drillhole L4D017 (left) and closed off RC hole L4R0215 (right)

All DD sites were rehabilitated by Günzel Drilling according to the site Environmental Management Plan, with foreign material removed and sumps filled and smoothed. Rehabilitation of the RC drillhole sites by Prinsloo drilling had only been partially completed at the time of the site visit as the drilling program was still in progress.

10.4. Areas 4 and 2B Reverse Circulation Drilling, 2023

A reverse circulation (RC) campaign was undertaken in 2023 at both Area 2B and Area 4 with the purpose of providing infill drilling data and to prove extension of the mineralisation both along strike and depth.

Drilling began in late January 2023, using an Atlas Copco CS 10 RC drill rig operated by Prinsloo Drilling, and continued until late November 2023. A total of 12 RC holes were collared at Area 2B, totaling 1 772 m of drilling, and 44 RC holes were completed at Area 4, totaling 9 043 m. Standard Operating Procedures were prepared by Gecko Exploration to guide the drilling and sampling activities.

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Source: Witley, 2023
Figure 57 RC Drilling at Area 4 (hole number L4R0218)

The majority of the 2023 drilling at Area 4 are infill holes which have resulted in a staggered 70 m drillhole spacing in combination with the previously completed diamond drillholes. The Area 2B RC drilling was collared predominantly in the northwest portion of the deposit at 100 spacing, resulting in an extension of the mineralisation along strike. A row of four holes were drilled in the southeast however these only targeted the shallower mineralised zones down to a depth of 100 m.

Collar locations for the completed drilling campaigns are shown in Figure 58 for Area 4 and Figure 59 for Area 2B. Down-hole surveys were conducted using a north seeking gyro (Boart Longyear True Gyro)

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Figure 58 Plan showing Area 4 drillhole collars

Figure 59 Plan showing Area 2B collars

10.5. Interpretation of Drilling Results

10.5.1. Area 4

Geological, lithogeochemical and geophysical surveys delineated an ENE-trending, REE-bearing alteration zone in Area 4. This zone was subsequently delineated by the 2011 drilling campaign to a depth of approximately 200 m. The mineralization is associated with a zone of variably intense albitisation and carbonatization that is centred along a major sinistral fault

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system. The 2011-2012 drilling campaign achieved a nominal drillhole spacing of 25 m for approximately 650 m along strike and demonstrated the down-dip continuity of the mineralised zone to vertical depths of more than 200 m. Four deep holes were drilled to test the continuity of the zone, and it was found to be present at vertical depths of up to 300 m.

As with the 2011 drilling, the 2012 infill holes were drilled perpendicular to the strike direction and were angled between 55° and 75° to intersect the southerly dipping mineralization at approximately right angles in an attempt to obtain near true thickness intersections. Eight holes (NLOFDH4084, NLOFDH4084B NLOFDH4085, NLOFDH4010, NLOFDH4011, NLOFDH4012, NLOFDH4013, and NLOFDH4014) were drilled down dip on the mineralization to recover sufficient drillhole core material for initial metallurgical test work. The positions of the drillhole collars are illustrated in Figure 54. Figure 60 shows an example of a typical drill section through the mineral deposit. The results of these drilling campaigns resulted in the declaration of an initial mineral resource in Area 4 (Siegfried and Hall, 2012).

The 2020 drilling campaign was planned to extend the mineral resource both along strike to the west and to greater depths. The 2012 drilling included four deep holes that indicated that the mineralization was continuous to a depth of up to 300 m vertically. However, these holes were not sufficiently closely spaced to be included in the previous mineral resource.

The recent drilling has extended the drilled strike length of the Area 4 altered/mineralized zone to approximately 1.5 km and has intersected the mineralized zone in multiple drillholes below 300 m, indicating that the mineralization is continuous to at least to this depth.

The orientation of the mineralized zone in 3D is well established by multiple drill intersections on close-spaced fences. Fence drilling indicated an orientation of between 070°E and 075°E and dips between 45°S and 60°S. This implies that the bulk of the drillholes intersected the targeted mineralization close to a 90° angle and that the difference between sample length and true thickness is therefore relatively minor.

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Source: Swinden, 2021
Figure 60 Example of a drilling section through the Main Zone mineralization at Area 4

Figure 60 illustrates that the Main Zone mineralization has variable grades of REE but consistently contains intervals with high Y concentrations (>0.1% Y).

Additional zones of REE mineralization with variable thickness occur up to 20 m to 40 m below the Main Zone and up to 25 m to 30 m above. These zones have potentially economic merit in an open-pit mining scenario.

10.5.2. Area 2B

The results of the diamond drilling at Area 2B, coupled with the results of geological mapping and trenching, show that the Area 2B zone represents a portion of a rare earth mineralizing system that is structurally controlled and was formed by both hydrothermal and magmatic (intrusive) events. Significant concentrations of REE (>0.5% TREE+Y) occur in narrow (< 1 m), generally carbonatized, zones of veining, fracture fill and breccia fill related to late hydrothermal activity within broader (10 m to 30 m) zones of albitic and carbonate alteration that are characterized by anomalous concentrations of HREE and Y.

Within the broader Area 2B alteration zone, the mineralization is shown by the drilling to occur in multiple, sub-parallel zones that have been traced from surface to vertical depths of about 200 m (Figure 61). Multiple intersections in drilling sections demonstrate that the alteration zones generally strike at approximately 045° to 055° and dip between 45° and 60° to the southeast. All drillholes were oriented approximately perpendicular to the mineralized zone in both plan and section. Intersections are considered to approximately reflect true width.

A key result of the 2020 drilling was to demonstrate that the zone previously indicated by 2010 drilling could be followed along strike between the widely separated holes. The mineralisation is now well defined along a strike length of slightly more than 800 m and to depths of at least 150 m.

Most holes in the Area 2B zone have a prominent shear zone at or near the footwall of the alteration zone. Most of the alteration occurs in the structural hanging wall. The alteration zone is of variable intensity and is variably mineralized along the Area 2B zone. Although much of the alteration sequence in the Area 2B zone has been enriched in REE, mineralization with

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significant grades appears to be relatively late in the alteration sequence. The specific mineralized structures are considered to be late veinlets, fracture fill, and breccia fill, and appear to be related to late hydrothermal activity following the main episode of albitization, carbonatization and related hydrothermal alteration.

Figure 61 Example of drilling section through the Area 2B mineralized zone

Note: Colour bands are ppm $Y_{2}O_{3}$. Outline of mineralization and alteration are interpreted from assay data.
Source: Swinden, 2021

10.6. Exploration Drilling Outside the Mineral Resource Areas

10.6.1. Location and Procedures

A total of 133 drillholes have been drilled on the Lofdal ML that were not part of the mineral resource drilling in Area 2B and Area 4. These holes were drilled between 2010 and 2020 on a variety of geological, lithogeochemical and radiometric targets. The location, orientation and length of these holes is given in

Table 22 and collar locations are shown in Figure 62. All holes were collared using HQ followed by NQ core where competent rock was intersected. These drillholes were drilled at the same time as the resource drilling campaigns in 2010, 2011, 2012, 2014 and 2020, using contractors, equipment and Standard Operating Procedures identical to those described for the mineral resource drilling. The comprehensive QAQC program described for the mineral resource drilling was implemented for all exploration drillholes.

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Note: Priority exploration areas are outlined by black lines. Resource areas (drillholes not shown) outlined in red lines.
Source: Background is Landsat Geocover Mosiac, 2000. Swinden, 2021
Figure 62 Location of exploration (non-resource) drillholes (white squares)

Table 22 Location and orientation information for exploration drillholes on the Lofdal ML. WGS84 UTM 33S

Hole ID	Area	Easting (m)	Northing (m)	Elevation (mamsl)	Azimuth (°)	Dip (°)	Depth (m)	End-Date
NLOFDH2021	Area 2	466 248	7 754 110	950.371	330	-55	155.4	2012-11-29
NLOFDH2022	Area 2	468 436	7 756 593	951.554	330	-55	62.3	2012-12-01
NLOFDH2023	Area 2	468 451	7 756 670	955.749	300	-55	59.5	2012-11-30
NLOFDH2024	Area 2	468 471	7 756 574	952.013	300	-55	83.3	2012-12-03
NLOFDH2025	Area 2	4691 41	7 757 594	933.55	320	-55	140.3	2012-12-06
NLOFDH2026	Area 2	469 176	7 757 625	935.109	320	-55	101.3	2012-12-07
NLOFDH2A014	Area 2A	467 890	7 754 970	947.524	10	-50	98.2	2010-10-22
NLOFDH2A015	Area 2A	467 934	7 754 952	948.043	14	-50	80.3	2010-10-25
NLOFDH2A016	Area 2A	468 033	7 754 959	947.326	8	-50	77.3	2010-10-27
NLOFDH2C017	Area 2C	467 385	7 754 369	964.916	339	-50	88.2	2010-10-30
NLOFDH2C018	Area 2C	467 067	7 754 192	986.665	337	-50	71.5	2010-10-28
NLOFDH2C019	Area 2C	467 085	7 754 235	988.507	295	-50	170.4	2010-10-31
NLOFDH2C020	Area 2C	466 959	7 754 140	976.524	336	-50	77.3	2010-11-03
NLOFDH4039	Area 4 NE Ext.	470 984	7 753 817	951.566	345	-55	110.3	2011-09-08
NLOFDH4040	Area 4 NE Ext.	471 736	7 754 103	959.319	345	-55	113.1	2011-09-09

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Hole ID	Area	Easting (m)	Northing (m)	Elevation (mamsl)	Azimuth (°)	Dip (°)	Depth (m)	End-Date
NLOFDH4041	Area 4 NE Ext.	471 023	7 753 704	952.819	345	-55	101.4	2011-09-10
NLOFDH4042	Area 4 NE Ext.	471 775	7 753 965	965.661	345	-55	101.3	2011-09-11
NLOFDH4043	Area 4 NE Ext.	471 050	7 753 574	950.024	345	-55	164.3	2011-09-12
NLOFDH4044	Area 4 NE Ext.	471 818	7 753 832	975.461	345	-55	151.1	2011-09-13
NLOFDH4045	Area 4 NE Ext.	471 028	7 753 654	952.216	345	-55	152.5	2011-09-14
NLOFDH4046	Area 4 NE Ext.	471 746	7 754 055	963.953	345	-55	152.1	2011-09-15
NLOFDH4101	Area 4 NE Ext.	472 008	7 754 265	952.729	330	-55	80.5	2012-09-07
NLOFDH4102	Area 4 NE Ext.	472 133	7 754 360	963.786	330	-55	68.3	2012-09-12
NLOFDH4103	Area 4 NE Ext.	472 263	7 754 426	973.797	330	-55	74.3	2012-09-11
NLOFDH4104	Area 4 NE Ext.	472 376	7 754 525	977.995	330	-55	101.3	2012-09-10
NLOFDH4105	Area 4 NE Ext.	472 399	7 754 487	974.818	330	-55	155.4	2012-09-14
NLOFDH4106	Area 4 NE Ext.	472 023	7 754 237	952.787	330	-55	82.2	2012-09-15
NLOFDH4107	Area 4 NE Ext.	472 158	7 754 317	956.442	330	-55	89.4	2012-09-18
NLOFDH4108	Area 4 NE Ext.	472 289	7 754 386	965.731	330	-55	113.4	2012-09-19
NLOFDH4109	Area 4 NE Ext.	472 724	7 754 747	976.635	330	-55	62.2	2012-09-21
NLOFDH5001	Area 5	468 785	7 754 614	956.973	305	-55	80.1	2011-06-08
NLOFDH5002	Area 5	468 783	7 754 586	957.114	305	-55	50.3	2011-06-09
NLOFDH5003	Area 5	468 712	7 754 572	960.006	305	-55	68.2	2011-06-10
NLOFDH5004	Area 5	468 673	7 754 540	962.062	305	-55	47.1	2011-06-11
NLOFDH5005	Area 5	468 738	7 754 556	959.106	305	-55	101.2	2011-06-12
NLOFDH5006	Area 5	468 576	7 754 390	967.638	305	-55	53.2	2011-06-13
NLOFDH5007	Area 5	468 602	7 754 380	967.279	305	-55	84.2	2011-06-14
NLOFDH5008	Area 5	468 460	7 754 197	967.535	305	-55	92.2	2011-06-15
NLOFDH5009	Area 5	468 481	7 754 189	968.04	305	-55	137.2	2011-06-16
NLOFDH5010	Area 5	468 430	7 754 162	970.155	305	-55	89.1	2011-06-17
NLOFDH5011	Area 5	468 451	7 754 143	970.337	305	-55	150.2	2011-06-21
NLOFDH5012	Area 5	468 400	7 754 118	972.429	305	-55	86.1	2011-06-22
NLOFDH5013	Area 5	468 426	7 754 104	972.276	305	-55	116.3	2011-06-23
NLOFDH5014	Area 5	468 195	7 753 748	983.625	305	-55	80.2	2011-06-24
NLOFDH5015	Area 5	468 216	7 753 732	985.07	305	-55	104.1	2011-07-05

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Hole ID	Area	Easting (m)	Northing (m)	Elevation (mamsl)	Azimuth (°)	Dip (°)	Depth (m)	End-Date
NLOFDH5016	Area 5	468 165	7 753 705	983.968	305	-55	77.3	2011-07-06
NLOFDH5017	Area 5	468 187	7 753 690	984.888	305	-55	110.3	2011-07-07
NLOFDH5018	Area 5	468 128	7 753 670	981.441	305	-55	77.3	2011-07-08
NLOFDH5019	Area 5	468 154	7 753 655	981.513	305	-55	89.3	2011-07-09
NLOFDH5020	Area 5	467 879	7 753 594	983.669	345	-55	104.3	2011-07-10
NLOFDH5021	Area 5	467 886	7 753 569	981.961	345	-55	128.3	2011-07-11
NLOFDH5022	Area 5	467 754	7 753 578	983.777	345	-55	53.4	2011-07-14
NLOFDH5023	Area 5	467 760	7 753 555	980.96	345	-55	77.4	2011-07-15
NLOFDH5024	Area 5	466 911	7 753 940	962.626	305	-55	50.3	2011-07-16
NLOFDH5025	Area 5	466 940	7 753 977	967.287	305	-55	62.4	2011-07-17
NLOFDH5026	Area 5	466 987	7 754 009	973.681	305	-55	71.4	2011-07-18
NLOFDH5027	Area 5	466 960	7 753 962	970.31	305	-55	86.2	2011-07-19
NLOFDH5028	Area 5	468 694	7 754 524	961.798	305	-55	89.3	2011-07-20
NLOFDH5029	Area 5	468 821	7 754 559	954.766	305	-55	92.3	2011-07-21
NLOFDH5030	Area 5	468 820	7 754 593	955.399	305	-55	98.3	2011-07-22
NLOFDH5031	Area 5	468 766	7 754 539	958.171	305	-55	137.4	2011-07-23
NLOFDH5032	Area 5	468 739	7 754 500	959.657	305	-55	140.3	2011-07-27
NLOFDH5033	Area 5	468 790	7 754 522	956.987	305	-55	140.3	2011-07-28
NLOFDH5034	Area 5	468 866	7 754 565	951.044	305	-55	161.2	2011-07-29
NLOFDH5035	Area 5	468 454	7 754 083	971.553	305	-55	191.3	2011-07-30
NLOFDH5036	Area 5	468 361	7 754 088	973.498	305	-55	77.2	2011-07-31
NLOFDH5037	Area 5	468 386	7 754 068	973.953	305	-55	122.4	2011-08-01
NLOFDH5038	Area 5	468 335	7 754 042	974.685	292	-55	92.3	2011-08-02
NLOFDH5039	Area 5	468 365	7 754 030	973.987	292	-55	122.3	2011-08-11
NLOFDH5040	Area 5	468 321	7 753 996	974.764	292	-55	89.3	2011-08-12
NLOFDH5041	Area 5	468 348	7 753 987	973.826	292	-55	131.5	2011-08-13
NLOFDH5042	Area 5	468 180	7 753 638	980.529	305	-55	161.9	2011-08-14
NLOFDH5043	Area 5	468 096	7 753 630	978.308	305	-55	80.5	2011-08-15
NLOFDH5044	Area 5	468 120	7 753 614	977.561	305	-55	110.3	2011-08-16
NLOFDH5045	Area 5	468 068	7 753 586	974.983	305	-55	71.2	2011-08-17
NLOFDH5046	Area 5	468 093	7 753 575	975.476	305	-55	116.3	2011-08-20
NLOFDH5047	Area 5	468 313	7 754 049	974.978	305	-55	80.4	2011-08-21
NLOFDH5048	Area 5	468 763	7 754 600	958.037	305	-55	50.2	2011-08-22
NLOFDH5049	Area 5	468 815	7 754 723	951.892	305	-55	53.2	2011-08-23
NLOFDH5050	Area 5	467 426	7 753 416	980.268	345	-55	77.3	2011-08-31
NLOFDH5051	Area 5	467 432	7 753 391	979.673	345	-55	107.4	2011-09-03
NLOFDH5052	Area 5	468 043	7 753 607	975.263	305	-55	149.1	2011-09-16
NLOFDH5053	Area 5	468 221	7 753 507	989.782	305	-55	110.0	2011-09-17
NLOFDH5054	Area 5	466 360	7 753 181	963.971	330	-55	152.3	2012-11-12

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Hole ID	Area	Easting (m)	Northing (m)	Elevation (mamsl)	Azimuth (°)	Dip (°)	Depth (m)	End-Date
NLOFDH5055	Area 5	466 583	7 753 215	953.333	330	-55	98.2	2012-11-14
NLOFDH5056	Area 5	466 546	7753 371	951.566	345	-55	140.1	2012-11-28
NLOFDH5057	Area 5	466 765	7 753 405	954.306	330	-55	215.4	2012-11-16
NLOFDH6001	Area 6	466 303	7 749 643	1 031.782	180	-55	122.3	2011-09-01
NLOFDH6002	Area 6	466 337	7 749 643	1 030.34	180	-55	122.4	2011-09-02
NLOFDH6003	Area 6	466 364	7 749 678	1 026.352	180	-55	122.3	2011-09-18
NLOFDH6004	Area 6	466 279	7 749 644	1031.453	180	-55	122.3	2011-09-19
NLOFDH6005	Area 6	465 759	7 749 415	967.483	330	-55	215.4	2012-09-24
NLOFDH6006	Area 6	465 696	7 749 558	981.915	150	-55	182.4	2012-09-28
NLOFDH6007	Area 6	465 688	7 749 575	982.414	150	-55	169.7	2012-10-04
NLOFDH6008	Area 6	465 583	7 749 507	970.518	150	-55	145.3	2012-10-18
NLOFDH6009	Area 6	465 567	7 749 535	976.727	150	-55	122.0	2012-10-20
NLOFDH6010	Area 6	466 136	7 750 078	1 026.085	150	-55	161.1	2012-10-25
NLOFDH6011	Area 6	466 227	7 750 368	1 053.167	150	-55	200.3	2013-06-04
NLOFDH6012	Area 6	466 440	7 749 718	1 021.493	150	-55	248.2	2013-05-04
NLOFDH6013	Area 6	466 193	7 749 906	1 035.132	150	-55	200.1	2013-05-08
NLOFDH6014	Area 6	466 233	7 750 084	1 028.034	150	-55	251.3	2013-05-12
NLOFDH6015	Area 6	466 027	7 750 080	1 010.091	150	-55	239.2	2013-05-29
NLOFDH6016	Area 6	466 147	7 750 282	1 044.728	330	-55	86.3	2013-05-21
NLOFDH6017	Area 6	466 056	7 750 469	1 050.965	150	-55	260.3	2013-05-25
NLOFDH6018	Area 6	466 300	7 750 465	1 034.338	150	-55	224.3	2013-06-08
NLOFDH6019	Area 6	465 812	7 749 682	1 019.716	150	-55	212.3	2013-06-14
NLOFDH6020	Area 6	465 856	7 749 210	973.838	150	-55	167.0	2013-06-17
NLOFDH6021	Area 6	465 445	7 749 436	979.531	150	-55	200.0	2013-06-20
NLOFDH6022	Area 6	466 301	7 749 974	1 026.312	150	-55	242.2	2013-07-06
NLOFDH6023B	Area 6	466 015	7749 814	1 026.714	150	-55	221.0	2013-07-12
NLOFDH6024	Area 6	465 954	7 749 914	1 013.825	150	-55	257.4	2013-07-19
NLOFDH7001	Main Intrusion	469 041	7 754 001	971.888	0	-90	239.6	2012-09-06
NLOFDH8001	Area 8	465 610	7 751 163	998.082	315	-55	152.3	2011-07-12
NLOFDH8002	Area 8	465 478	7 751 319	988.551	135	-55	152.3	2011-07-13
NLOFDH8003	Area 8	465 708	7 751 470	981.675	0	-90	152.5	2011-07-24
NLOFDH8004	Area 8	465 246	7 751 083	990.618	0	-55	80.2	2011-07-25
NLOFDH8005C	Area 8	465 455	7 751 355	983.709	324	-55	182.1	2012-11-05
NLOFDH8006	Area 8	465 516	7 751 270	995.812	324	-55	149.9	2012-11-07
NLOFDH8007	Area 8	465 651	7 751 102	999.671	324	-55	152.0	2012-11-09
LDD0001	Dolomite Hill	468 829	7 755 507	974.266	310	-60	65.6	2020-07-30
LDD0002	Dolomite Hill	468 874	7 755 548	966.517	330	-80	53.9	2020-08-01

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Hole ID	Area	Easting (m)	Northing (m)	Elevation (mamsl)	Azimuth (°)	Dip (°)	Depth (m)	End-Date
LDD0003	Dolomite Hill	469 121	7 755 809	947.552	310	-55	104.4	2020-08-04
LDD0004	Dolomite Hill	469 149	7 755 802	950.501	310	-55	152.9	2020-08-06
LND0001	North Splay	478 726	7 758 238	943.392	330	-55	77.7	2020-06-26
LND0002	North Splay	478 766	7 758 268	938.433	330	-55	68.7	2020-06-27
LND0003	North Splay	478 777	7 758 247	939.553	330	-55	122.5	2020-07-08
LND0004	North Splay	478 814	7 758 295	934.592	330	-55	83.7	2020-07-10
LND0005	North Splay	478 715	7 758 163	937.839	330	-55	215.9	2020-07-14
LND0006	North Splay	478 630	7 758 163	934.675	330	-55	116.8	2020-07-16
LND0007	North Splay	478 409	7 758 063	932.468	330	-55	86.7	2020-07-18
LND0008	North Splay	478 417	7 758 037	931.765	330	-55	143.8	2020-07-21
LND0009	North Splay	478 285	7 757 971	938.997	330	-55	140.9	2020-07-23
LND0010	North Splay	478 297	7 757 940	933.545	330	-55	218.9	2020-07-28

10.6.2. Exploration Drilling Results

None of the exploration drilling outside of Area 2B and Area 4 has identified mineralization of a size and grade that it could be included as part of the mineral resource, and these exploration drilling results are not considered to be material to the mineral resource on the Lofdal ML. Results are referenced to exploration priority areas shown on Figure 62. Assays of typical significant mineralized intersections are given in Table 23.

10.6.2.1. Area 2

Exploration holes were drilled in Areas 2A and 2C as part of the initial drilling campaign that defined the alteration/mineralization zone in Area 2B. These holes targeted narrow exposed carbonatite dykes that yielded individual grab samples exhibiting high grades of HREE. The objective of this drilling was to test different styles and grades of dykes as a prelude to a more comprehensive drilling program in subsequent years. The drilling demonstrated that the narrow dykes exposed at surface in Area 2 did not widen with depth and anomalous mineralization was only present over widths of less than 2 m.

In 2012, six drillholes were drilled to test the northeastern extension of the Area 2B alteration system approximately 2 km to 3 km NE of Area 2B. These holes targeted the alteration/mineralization zone beneath where the best surface grab samples were taken from this part of the zone where albitic and carbonatitic alteration were well developed. Three of these holes intersected narrow widths (3 m to 4 m) of relatively low-grade mineralization, demonstrating that although the zone was carrying REE in this area, it is relatively low grade and discontinuous.

10.6.2.2. Area 4 NE Extension

The alteration zone that hosts the Area 4 mineral resource can be traced along strike in outcrop to the northeast for almost 9 km. Surface grab sampling outlined a number of areas with anomalous REE contents (described in Dodd et al., 2014). Seventeen drillholes were drilled in this zone in 2011 and 2012 to test the width and grade of the zone within approximately 2 km of the east end of the Area 4 mineral resource. Although the alteration was intersected in most holes, REE values were uniformly low grade, and anomalous over less than 4 m of core length.

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10.6.2.3. Area 5

Area 5 is an extensive alteration zone with associated REE and thorite mineralization that outcrops along the northwest side of the Main Intrusion. Outcrops at its northeast end returned good REE values and the zone returned anomalous surface grab samples along more than 3.3 km of strike length. Targets were defined by surface outcrops with anomalous grab samples and by radiometric anomalies. 53 drillholes were drilled to test this zone. Mineralization was encountered in most drillholes, but it was found to be inconsistent in grade and width. The best intersections were between 10 m and 15 m wide (not true width; the orientation of the zone is not well defined by the drilling) with moderate grades of REE. However, significant intersections could not be connected between holes on the same section or along strike.

10.6.2.4. Area 6

Area 6 is an aerially extensive zone of fenetization associated with an intense radiometric anomaly in sedimentary rocks immediately north of the Oas Syenite. Surface outcrops are generally not well mineralized and where present, form highly radioactive, very narrow veins. 24 holes were drilled in this area in an attempt to define the nature and extent of any REE mineralization. The drilling found extensive zones of relatively low grade REE over a considerable vertical extent, but no areas were identified with significant grades that could be connected by several drillholes or that would be considered of economic interest. The alteration and mineralization do not seem to occupy consistent structures and there is no indication from the drilling as to whether the mineralized widths are true widths. The alteration contrasts in mineralogy and lithology with other mineralized alteration zones at Lofdal. The REE mineralization is also different, dominated by REE silicates (britholite, allanite), rather than phosphates, and typically associated with fluorite and locally molybdenite.

10.6.2.5. Area 8

Area 8 comprises the Emanya Plug and a number of nearby carbonatite dykes. There was little indication of mineralization in this area from surface grab samples or radiometrics. Seven drillholes were completed to test the potential of this plug and only very low grade, sporadic mineralization was encountered.

10.6.2.6. Dolomite Hill

Dolomite Hill is a wide zone of alteration immediately north of the Main Intrusion that returned some relatively high grades from surface grab sampling. Four holes were drilled in 2020 to test whether the size and grade of this zone shows improvement with depth. The holes encountered only sporadic mineralization, with few consecutive samples exhibiting anomalous REE mineralisation.

10.6.2.7. North Splay

The North Splay is an outcropping albitite and carbonatite alteration system. It is the most northerly and the most distant mineralization from the Main Intrusion and Area 4. A number of anomalous grade grab surface samples resulted from sampling that occur over a strike length of approximately $1.4\mathrm{km}$, as described by Dodd et al. (2014). Despite the fact that some alteration was encountered in the core, there were no REE values of potential economic interest in the core samples.

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Table 23 Analyses of typical significant altered/mineralized intersections in exploration drillholes

Hole ID	From m	To m	width m	La ppm	Ce ppm	Pr ppm	Nd ppm	Sm ppm	Eu ppm	Gd ppm	Tb ppm	Dy ppm	Ho ppm	Er ppm	Tm ppm	Yb ppm	Lu ppm	Y ppm	THRE E+Y %	TREE +Y %
NLOFDH2A014	38	39	1	493	971	97	381	109	44	158	28	167	29	83	11	75	11	1 045	0.16	0.36
NLOFDH2021	105	108	3	193	366	39	160	106	47	173	35	215	37	96	13	78	11	978	0.16	0.25
NLOFDH4104	35	38	4	73	137	15	63	66	29	121	21	111	18	41	5	28	4	437	0.08	0.11
NLOFDH5051	51	62	11	899	1 543	163	582	166	55	190	27	135	23	62	9	57	9	616	0.20	0.54
NLOFDH5012	46	61	15	57	117	13	57	92	53	222	43	254	48	125	17	99	14	1 258	0.27	0.31
NLOFDH6008	87	94	7	596	883	83	282	96	33	116	19	113	21	59	8	47	6	632	0.24	0.44
NLOFDH8005C	129	130	6	6 400	10 235	948	2 776	263	61	154	18	94	17	43	6	31	4	565	0.26	2.32

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11. Sample Preparation, Analyses and Security

Sample preparation for the 2010 and 2011-2012 drilling campaigns and for lithogeochemical grab sampling has previously been described in detail by Swinden and Siegfried (2011) and Siegfried and Hall (2012). Each drilling campaign had its own set of Standard Operating Procedures. These were similar to those employed in the 2020 drilling and were implemented in accordance with the CIM Best Practice Exploration Guidelines (Refer to Table 19 and Table 21 for summaries of procedures during these drilling campaigns).

The standard operating procedures (SOPs) for geological and geotechnical logging, core splitting and sampling were compiled by Gecko and reviewed by MSA to ensure that the various activities were carried out in a consistent, transparent, auditable and appropriate manner in accordance with industry standards.

11.1. Diamond Drilling Procedures

The following descriptions refer to procedures followed during the 2020 drilling.

11.1.1. Drillhole Logging

Geological and geophysical logging (Gamma logging of all core; handheld PXRF for Y on core with anomalous radiometric readings) was carried out by Gecko geo-technicians and geologists and followed a comprehensive protocol.

Structural data, alpha and beta angles (Figure 63), were collected on the core to determine the spatial orientation of mineralising and barren vein systems. The alpha angle is the acute angle between the core axis and the long axis of the ellipse (0°-90°; Figure 63). The beta angle is the angle between the orientation line marking "Top of Hole" as reference line along the core and the ellipse apical trace measured in a clockwise sense (0°-360°; Figure 63). Alpha and beta angles data were collected using a goniometer in all mineralised zones and other zones as deemed necessary by the logging geologist.

Source: Holcombe, 2016
Figure 63 Illustration of the alpha, beta, and gamma angles in core

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The drillhole cores were logged in detail, recording lithology, alteration, intense near surface weathering / overburden and structure. The weathering at depth is visually indistinct but is evident from core recovery and geotechnical logging (RQD).

11.1.2. Sample Preparation

As mineralization is not always visually discernible, the core intervals to be sampled were determined at the discretion of the logging and sampling geologist using both logging information as well as gamma readings measured with a RadEye PRD scintillometer, and XRF analyses obtained from an Olympus Delta 50 or Olympus Vanta handheld XRF analyser (PXRF). The RadEye was used to obtain indicative gamma readings along the entire length of core while in the box and the maximum and minimum values were recorded. Where the in-box gamma values were greater than 50 the core was removed from the box and the RadEye was used to determine gamma values again. PXRF readings for Y were taken along the entire core and the values noted. Both the RadEye gamma and PXRF Y values are indicative and are used solely for the purpose of identifying areas to be sampled for laboratory analysis.

Sampling of the drillhole core was undertaken after metre marking, geological and geotechnical logging, and photographing of the core. All core cutting, sampling, bagging and dispatch procedures were undertaken at the Lofdal field camp. After this work was completed, the remaining core was transported to the warehouse in Khorixas for storage.

Mineralised intervals in the drillhole core were generally sampled at one metre intervals. In cases where lithological changes were observed within a one metre sampling interval, then each lithology was sampled separately, using a minimum 15 cm core length. Sampling was to at least 2 m above and below the zone identified as potentially mineralised. Narrower potentially mineralised zones away from the main zones of mineralisation were also sampled and, at the discretion of the sampling geologist, a single one metre sample was taken either side of the potentially mineralised zone.

11.1.2.1. Core Marking and Splitting

Prior to cutting the core, sampling intervals and unique sample numbers (sample ticket book number) were clearly marked above the core orientation line and below the core cutting line drawn on the drillhole core. In instances where the orientation of the core was unknown, then the cutting line and the orientation line were the same line. The start and end of each sample was marked with a yellow line around the core and a white dot on the core cutting line.

A designated geologist responsible for all core sampling carried out the core sampling. The colour convention of yellow (metre mark), white (sampling interval) and red (sample number) was used for all drillholes. (Figure 64).

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Source: Ellmies, 2020
Figure 64 Examples of drillhole core marking

The core was split in half using a commercial core cutter with a 2.2 mm wide diamond core cutting blade (Figure 65). The split halves were returned to the labelled core boxes between the depth blocks and correctly orientated with the aid of the downhole arrows. The logging/sampling geologist checked the core boxes to ensure that all core and core markings were correct prior to removing the sample. The upper half of the core was used for analysis and the lower half of the core was retained in the core tray for future reference or additional test work. Sample numbers were marked on each individual piece of core with a red waterproof marker and recorded in the customised sampling sheet, which was then captured in the project database.

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Source: Ellmies, 2020
Figure 65 Core cutting device

11.1.2.2. Core Sampling and Sample Dispatch

Each core sample was assigned a unique number on a wet strength, sequential sample number tag and the sample, generally representing one metre, was consistently taken from the same side of the core relative to the cutting line and placed in a thick plastic sample bag. Two sample number tags were placed in each sample bag, one inside the bag and the other clearly visible to the outside of the bag but located in the folded seal of the bag. Bags were securely sealed with staples and sequentially packed ready for dispatch. Drillhole number and sampling interval were recorded on the sampling book stub and entered into customized sampling sheets which were then digitally captured into the on-site computer. The sampling database was regularly transmitted and backed up at NMI's Windhoek office.

Gaps in the sample sequence were left for standards, blanks and duplicates during the sampling process. The standards and blanks were only packed and labelled with the assigned sample numbers after the core sampling process was completed, in order to minimise the possibility that sample numbers are inadvertently swapped between routine samples, standards or blanks.

The geologist responsible for sampling and dispatch verified the sample numbers and sequence before the samples were packaged in groups of ten into uniquely numbered heavy-duty bags, which were closed with cable ties. The bags were then re-checked against the final sample submission sheet and signed off by the geologist before being loaded and transported in a company vehicle to Actlabs Namibia (Pty) Ltd (Actlabs) in Windhoek.

The Gecko driver of the vehicle signed two copies of an acceptance/transportation sheet specifying the quantity of bags together with sample export documentation for the onward dispatch to Actlabs in Canada. The samples were dispatched to Windhoek on a weekly or fortnightly basis and all sample transport documentation is filed at the NMI offices in Windhoek.

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11.1.2.3. Density Measurements

Rock density measurements using the Archimedes principle (weight in air versus weight in water) were taken for a portion of each core sample, after splitting and sampling. Each density sample was between approximately 15 cm and 20 cm long. The density device comprises a 3 kg electronic scale, below which a water container was placed. A core sample holder attached to the balance was used to immerse core in water in the container. The method was as follows:

the balance was reset to 0.00 g before each reading;
a dry length of core was placed in the core holder and the mass of the core in air was recorded;
the container was filled with water to submerge the sample and the mass of the core was determined in water.

The density (specific gravity or SG) was calculated using the formula:

$$
SG = \frac{\text{Mass in Air}}{\text{Mass in Air} - \text{Mass in Water}} = \frac{W}{V}
$$

Source: Ellmies, 2020
Figure 66 Apparatus for measuring density (SG) by the Archimedes principle

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11.1.2.4. Core Storage

The core trays with the unsampled intervals and remaining halves of the sampled intervals are permanently stored in a rented facility near Khorixas (Figure 67). The fenced premises are locked, as is the warehouse and all storage containers. Only Gecko and NMI staff and NMI consultants have access to the building where the core is stored.

Source: Ellmies, 2020

Figure 67 Core storage in the Khorixas warehouse and in containers in the warehouse yard

11.2. Reverse Circulation Drilling Procedures

Prior to drilling, three bags were pre-labelled with a permanent marker. Drilling was undertaken at one metre intervals resulting in material consisting of rock chips and dust. These were transported to surface and channelled into a rig-mounted, three-tier riffle splitter. The riffle splitter generated an A-sample and a B-sample, weighing between 1.8 kg and 4.5 kg, and the remainder as a bulk sample with an average weight of 28 kg. Samples were weighed immediately after collection and secured with cable ties. The A-sample / B-sample bags were placed in larger polyweave bags, each containing ten sample bags.

11.2.1. Reverse Circulation Logging

A small portion of the sample was passed through a 50 µm aperture sieve to separate the coarse chips from the fines, the coarse chips were washed and collected in pre-labelled chip trays for logging purposes.

The mineralised zone was identified through gamma readings of the core chips using a "Radeye" Geiger counter by Thermo Fisher Scientific GmbH. The Radeye was placed on top of the sample bag for each drilled metre for a period of ten seconds. The average number of counts per second (cps) was recorded by the geologist and if an average of more than 50 cps was recorded, the sample bag was placed one metre away from the adjacent bags and re-analysed. This was done to confirm the gamma readings and minimise the influence from the adjacent bags. Once the mineralised zone was defined, geologists were instructed to visually re-log the chips from this zone.

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The logging recorded primary and secondary lithologies, alteration, colour and visible minerals. These were captured on an NMI logging template.

The 50 µm RC fines for each drilled metre were analysed using a handheld Olympus Inno-X Delta Dynamic X-Ray Fluorescence analyser. All samples were analysed twice using the "Mining" mode for an average of 10 to 20 seconds. Zones containing samples with yttrium values greater than 100 ppm, together with the samples identified via gamma readings were selected for submission for laboratory analysis.

11.2.2. Chip Sampling and Sample Dispatch

Sampling lists, with Quality Assurance and Quality Control (QAQC) samples were pre-populated, prior to commencement of drilling. Sampling was only undertaken once the mineralised zone was defined using the XRF and gamma logging. The A-sample derived from the riffle splitter was used for submission to the laboratory, sample tickets were stapled to these bags and QAQC samples were inserted in their pre-defined sequence. The drillhole ID number and the sample depths were removed from the bags using paint thinner. Ten samples were placed inside a polyweave bag and sealed. Any A-samples not submitted to the laboratory were stored at the NMI warehouse in Khorixas, together with the B samples.

Samples were submitted to the laboratory in batches, consisting of two drillholes per batch. The samples were transported to Actalabs Namibia by NMI or Gecko staff.

11.3. Sample Analyses

Samples for the diamond drilling and RC campaigns underwent the same sample preparation and analytical procedure.

11.3.1. Sample Preparation at the Laboratory

At the Actlabs preparation facility in Windhoek, the samples were laid out and checked against the NMI sample list to verify that all samples are present and correctly numbered. An internal sample tracking sheet was prepared by Actlabs to track progress of the samples through the laboratory.

Using Actlabs' sample preparation package RX1, the samples were initially crushed in a jaw crusher to 80% passing two mm and then passed through a riffle splitter to obtain a 250 g split for pulverisation. The splits were pulverized with a swing mill in hardened steel bowls to 95% passing 105 µm. Samples were then homogenized in a stainless-steel riffle splitter and a 15 g sample and duplicate were drawn from the splitter for analysis. The splits were placed in Ziploc bags and prepared for shipping to Actlabs' analytical laboratory in Canada. The duplicate pulps were stored at the Actlabs facility in Windhoek.

11.3.2. Sample Analyses at the Laboratory

The pulp samples were couriered by air to the Actlabs analytical facility in Ancaster, Ontario, Canada, where they were analysed for major element oxides, rare earth elements and other trace elements.

Actlabs used their Code 8, REE Assay Package which involves a lithium metaborate fusion, multi acid digestion, and Inductively Coupled Plasma Analysis – Optical Emission Spectroscopy (ICP-OES) finish for major element oxides; Sc, Be, V, Sr, Y, and Zr. An ICP-Mass Spectrometry (ICP-MS) finish was used for other trace elements including the REE. Nb₂O₅ and ZrO₂ were

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determined by sample fusion and standard X-ray fluorescence (XRF) method for samples with >0.3% P₂O₅.

Rare earth elements are among the most difficult elements to analyse to a high degree of analytical precision under a wide range of individual REE concentrations. The lithium metaborate fusion and ICP-MS finish is the current industry standard for high quality REE analyses.

Actlabs' quality system is accredited to international quality standards through the International Organisation for Standardisation / International Electrotechnical Commission (ISO/IEC) 17025, which includes ISO 9001 and ISO 9002 specifications, with CAN-P-1758 (Forensics), CAN-P-1579 (Mineral Analysis) and CAN-P-1585 (Environmental) for specific registered tests by the Standards Council of Canada (SCC).

11.4. Sample Security

All drillhole core handling, sampling and transportation activities were undertaken by Gecko staff. The individual procedures followed strict protocols outlined in a comprehensive SOP manual which was drafted by Gecko and reviewed by MSA. NMI's field camp has sample preparation facilities and is located in a relatively remote area to which only staff and contractors have access.

The core boxes were transported from the drilling rigs to the exploration camp on a daily basis by Gecko staff. To reduce movement of the core, the boxes were covered with foam sheets and ratchet-strapped to the loading bay of a utility vehicle for transport. Once the samples had been taken and prepared for dispatch, a Gecko staff member transported the samples in sealed bags to Actlabs' Windhoek preparation laboratory from where the material was couriered to Actlabs' Canadian facilities for sample analyses.

A "chain of custody" is maintained from the site to the laboratory via locked facilities and dispatch and receipt documentation. MSA considers that there was little or no opportunity for sample tampering by an outside agent due to the secure and auditable "chain of custody" implemented by Gecko and NMI personnel.

11.5. Quality Assurance and Quality Control

Appropriate quality assurance and quality control (QAQC) monitoring is a critical aspect of the sampling and assaying process in any exploration program. Monitoring the quality of laboratory analyses is fundamental in ensuring the highest degree of confidence in the analytical data and providing the necessary confidence to make informed decisions when interpreting all the available information. QA may be defined as information collected to demonstrate that the data used in the project are valid. QC comprises procedures designed to maintain a desired level of quality in the assay database. Effectively applied, QC leads to identification and correction of errors or changes in procedures that improve overall data quality. Appropriate documentation of QC measures and regular scrutiny of QC data are important as a safeguard for project data and form the basis for the quality assurance program implemented during exploration.

In order to ensure quality standards are met and maintained, planning and implementation of a range of external quality control measures is required. Such measures are essential for minimizing uncertainty and improving the integrity of the assay database and are aimed to provide:

An integrity check on the reliability of the data;
Quantification of accuracy and precision;
Confidence in the sample and assay database;
The necessary documentation to support database validation.

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For all its drilling campaigns at Lofdal, NMI has adopted an industry standard QAQC program and inserted internal standards and certified reference material (CRM) and blanks each at a frequency of one in 20 (5%) into the batches prior to submission to Actlabs. These control samples were inserted as part of a continuous sample number sequence and the QAQC samples were not obviously different from routine samples after the pulverization process. Actlabs were requested in the sample submission sheet to split the pulp of predetermined samples (1 in 20) and insert the material in the empty and pre-numbered bags to create the required 5% duplicate samples. Actlabs in Canada was unaware which samples were QAQC samples and what their composition was. This allowed for monitoring of the sample preparation procedure as well as monitoring the accuracy and precision of analyses.

5% of the sample assayed by Actlabs were submitted to a second laboratory in Canada for check analysis. Hence the overall number of control samples constituted 20% of all samples analysed, which is in line with good practice procedures to ensure integrity of data and is independent from the internal QAQC methods applied by the laboratory itself.

11.5.1. 2010, 2011 and 2012 Drilling Programme

Results of the 2010, 2011 and 2012 QAQC program were reported by Swinden and Seigfried (2011) and Seigfried and Hall (2012). The QAQC programs for these campaigns demonstrated that the analytical work was fully adequate and did not indicate any issues with the data quality.

11.5.2. 2020 Drilling Programme

Data from duplicates, internal standards, CRMs and blanks were examined on a batch-by-batch basis to immediately identify any errors in the analytical data. Data from duplicate, standard and blank analyses was examined numerically and graphically to determine the repeatability of the duplicate analyses, the accuracy of the standard analyses with respect to the accepted values, and the levels of REE present in the blanks.

11.5.2.1. Blank Samples

Two different blank materials were used to evaluate sample preparation:

Quartz pebble blank sourced from Ferreiras Garden Shop in Windhoek. Used in drillholes L4D0115 through L4D00136 and in drillholes L2BD0027 through L2BD0041; and
Dolomite sourced from Ferreiras Garden Shop in Windhoek. Used in drillholes L4D0137 through L4D00170 and in drillholes L2BD0042 through L2BD0055.

The blank materials were supplied as coarse gravel and a sample of approximately 50 g was used. The blank samples were inserted with consecutive numbers within the core sample stream and underwent the same sample preparation and analytical processes as the routine field samples. Graphical representations of blank sample results for selected REE are shown in and Figure 69 and blank sample summaries for all REEs are given in

Table 24. Repeated analyses show that these materials are acceptable for use as blanks for carbonatite hosted REE analyses with all REE present in concentrations near or below their detection limits. Background values are slightly higher in the quartz pebble blank and this is reflected by a number of slightly anomalous and insignificant (i.e., between 1 and 3 ppm) Ce analyses in this material. There are no anomalies in the analyses that suggest anything more than normal difficulties of analysing these elements at very low concentrations. No further action was taken or required, and the results of the blank analyses indicate that there was no

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contamination or systematic analytical issues during the period of sample submission and analyses.

Note: vertical axis in ppm
Figure 68 Blank analyses for selected REE from Area 2B

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Note: vertical axis in ppm
Figure 69 Blank analyses for selected REE from Area 4

Table 24 Number of blank failures (>10 times LDL*)

Number of Samples	La	Ce	Pr	Nd	Sm	Eu	Gd	Tb	Dy	Ho	Er	Tm	Yb	Lu	Y
300	6	33	1	2	0	0	0	0	0	0	0	0	0	0	0

Note: *LDL=Lowest Detection Limit

11.5.2.2. Standards

During the 2020 drilling campaign, four different standards were used. Three NMI in-house standards with varying degrees of HREE enrichment, namely STD4, STD5 and STD6, were used. All in-house standards underwent 'round robin' analyses at four independent laboratories but are not certified. Repeated analyses of STD4 and STD6 throughout the program demonstrated a generally consistent REE composition. The use of STD5, however, was discontinued after 20 holes because of reproducibility issues, particularly with respect to Sm, Tb, Er and Tm. One commercial CRM, AMIS0185 was also used. The heavy rare earth elements Gd, Tb, Er, Yb and Lu in this CRM are neither certified nor provisional but only reported for informational purposes and were not used for the purposes of QAQC in this project. Control charts showing the assays through time of several REEs for the 2020 drilling are shown in Figure 70 to Figure 76 and the standards data for all elements is summarized in Table 25 and Table 26.

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Figure 70 Analyses of the CRM AMIS0185 for selected REE in Area 2B

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Figure 71 Analyses of the CRM AMIS0185 for selected REE in Area 4

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Figure 72 Analyses of the Standard STD4 for selected REE in Area 2B

Figure 73 Analyses of the Standard STD4 for selected REE in Area 4

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Figure 74 Analyses of the Standard STD5 for selected REE in Area 4

Figure 75 Analyses of the Standard STD6 for selected REE in Area 2B

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Figure 76 Analyses of the Standard STD6 for selected REE in Area 4

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Table 25 Statistics for the reference materials used in the 2020 drilling program

CRM Name	Number Used	Statistic	La	Ce	Pr	Nd	Sm	Eu	Gd	Tb	Dy	Ho	Er	Tm	Yb	Lu	Y
STD 4	81	Accepted Mean	622.0	990.2	102.0	398.2	179.9	76.6	300.6	16.1	396.3	81.1	227.8	33.7	210.8	30.0	2 439.7
		Standard Deviation	36.9	5.6	22.6	4.6	3.3	14.0	2.1	16.1	10.6	4.4	6.6	0.9	6.5	2.2	97.9
STD 5	18	Accepted Mean	1 089.3	1 666.2	173.9	739.2	465.1	201.4	787.6	39.4	785.9	142.8	370.1	49.7	276.0	36.8	4 126.0
		Standard Deviation	65.3	7.9	39.3	8.2	9.7	25.5	3.0	39.4	23.3	6.9	5.5	0.6	9.2	2.4	162.6
STD 6	88	Accepted Mean	175.7	331.7	38.8	161.9	178.5	137.1	804.4	6.2	1 758.9	399.2	1 228.2	183.7	1 106.7	146.4	12 659.0
		Standard Deviation	11.9	1.2	7.2	7.6	5.5	38.1	8.4	6.2	51.9	12.6	63.1	4.9	34.8	10.9	468.0
AMIS0185*	106	Accepted Mean	29 760.0	40 750.0	3 471.0	9 238.0	556.0	94.2	29 760.0	2 720.0	27.1	3.2	-	0.4	-	-	62.0
		Standard Deviation	4 610.0	343.00	1033.0	48.00	12.10	2 720.0	-	-	5.10	0.50	-	0.08	-	-	7.70

Note: * Certified and provisional concentration only

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Table 26 Failure rate (outside ± 3 SD) for standards assayed by Actlabs during the 2020 drilling campaign

CRM Name	Failure Rate	La	Ce	Pr	Nd	Sm	Eu	Gd	Tb	Dy	Ho	Er	Tm	Yb	Lu	Y
STD 4	Number of Samples	7	3	3	3	17	3	2	4	8	2	6	7	7	2	2
	Percentage	6.5 %	2.8 %	2.8 %	2.8 %	15.7 %	2.8 %	1.9 %	3.7 %	7.4 %	1.9 %	5.6 %	6.5 %	6.5 %	1.9 %	1.9 %
STD 5	Number of Samples	1	1	1	1	8	1	0	4	1	1	6	9	1	1	1
	Percentage	5.6 %	5.6 %	5.6 %	5.6 %	44.4 %	5.6 %	0.0 %	22.2 %	5.6 %	5.6 %	33.3 %	50.0 %	5.6 %	5.6 %	5.6 %
STD 6	Number of Samples	1	2	4	0	0	0	0	3	5	1	0	10	3	0	0
	Percentage	1.1 %	2.3 %	4.5 %	0.0 %	0.0 %	0.0 %	0.0 %	3.4 %	5.7 %	1.1 %	0.0 %	11.4 %	3.4 %	0.0 %	0.0 %
AMIS0185*	Number of Samples	0	0	0	0	0	0	-	-	0	0	-	-	-	-	1
	Percentage	0.0 %	0.0 %	0.0 %	0.0 %	0.0 %	0.0 %	-	-	0.0 %	0.0 %	-	-	-	-	0.9 %

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Four samples gave highly anomalous results and required further investigation as detailed in Table 27. Three were found to result from incorrect categorisation of the sample material, probably taking place in the core yard, and the other was due to a missing Y assay. None indicate an issue with the sample preparation or analytical procedures.

Table 27 Resolution of anomalous CRM analyses
| Area | Standard as recorded | Elements affected | Resolution |
| --- | --- | --- | --- |
| Area 4 | STD 4 | All REE display a positive spike | STD5 was mistakenly inserted |
| Area 4 | STD5 | Negative spike in LREE, positive spike in HREE | STD6 was mistakenly inserted |
| Area 4 | AMIS 185 | Zero value for one Y analysis | Cell was blank on the reporting spreadsheets |
| Area 2B | STD4 | All REE display a negative spike | A blank was mistakenly inserted |

11.5.2.3. Pulp Duplicates

Laboratory duplicates were prepared for every 1 in 20 samples. The original and duplicate analyses were compared graphically to ensure repeatability. Where significant outliers from the expected values were observed in adjacent control samples, a subset of the batch was re-analysed. This was the case in one batch where several duplicates displayed an anomalous amount of divergence. Re-analyses produced acceptable duplicate analyses. No additional measures were taken or deemed necessary.

Note: Reference lines are: 1:1; +10% and -10%, values are ppm
Figure 77 Analyses of laboratory duplicates for selected REE from Area 2B

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Figure 78 Analyses of laboratory duplicates for selected REE from Area 4

Note: Reference lines are: 1:1; +10% and -10%, values are ppm

Table 28 Percentage of assays within mean absolute difference of 10% and 20% (above 10x LDL) – Actlabs duplicate versus original

	La	Ce	Pr	Nd	Sm	Eu	Gd	Tb	Dy	Ho	Er	Tm	Yb	Lu	Y
10%	76	77	78	82	77	80	80	79	82	81	80	79	82	80	83
20%	94	95	95	96	97	97	97	97	96	96	97	96	96	96	96

11.5.2.4. Second Laboratory Duplicate Assays (“Umpire Laboratory”)

Approximately 5% of samples were sent for check analyses at a second laboratory; ALS Minerals (ALS) in North Vancouver, Canada. Pulps were split from analytical samples at the Actlabs sample preparation facility in Windhoek and were shipped directly to ALS. The results of these analyses were plotted graphically against the original analysis. In the vast majority of duplicate sample pairs, there was less than 10% difference. The results for selected REEs are presented in Figure 79 and summarised for all REEs in Table 29.

ALS are registered to ISO 9001:2008 and have received ISO 17025 accreditation for laboratory procedures relevant for the purpose of the check assay exercise.

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Figure 79 Scatterplot of duplicate pair data (Actlabs and the umpire lab (ALS)) for selected REES

Note: ALS assays on horizontal axis. Actlabs on vertical axis, values in ppm
Blue reference lines is 1:1; red lines are +10% and -10%

Table 29 Mean and Variance of original and duplicate data – Actlabs versus ALS

Element	Actlabs		ALS
	Mean	Variance	Mean	Variance
La	118.6	62 340.2	117.0	67 455.5
Ce	221.7	203 810.4	216.7	204 241.6
Pr	24.5	2 609.9	24.1	2 611.4
Nd	99.3	70 015.8	96.1	63 104.1
Sm	31.9	11 530.6	30.8	10 166.8
Eu	10.5	820.0	10.6	850.5
Gd	41.1	9 203.8	38.6	7 464.5
Tb	7.2	210.3	7.2	207.9
Dy	44.9	9 701.8	45.4	9 591.9
Ho	9.2	455.8	9.0	471.4
Er	27.5	5 527.6	26.2	4 656.1
Tm	3.9	108.0	3.8	113.1
Yb	25.1	5 379.7	24.0	4 765.2
Lu	3.3	80.8	3.5	106.2
Y	278.5	635 284.9	267.6	509 003.4

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11.5.3. 2023 RC Programme

The QAQC for the 2023 RC drilling programme consisted of blanks, field duplicates, lab duplicates and CRMs with an average insertion rate of 1 in every 10 samples.

11.5.3.1. Blanks

A total of 284 blanks were used for the 2023 RC programme which underwent the same sample preparation and analytical process as the primary samples. MSA checked the results for the blanks samples for individual REE elements using a threshold of 10 times the lower detection limit. No significant levels of contamination were detected for the HREEs, however several LREEs reported values above the threshold of 1 ppm, with La values up to 7.4 ppm and Ce values up to 10 ppm. The blank chart for dysprosium is shown in Figure 80 and for cerium in Figure 81.

Figure 80 2023 RC programme blank chart for dysprosium

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Figure 81 2023 RC programme blank chart for cerium

11.5.3.2. Standards

The same standards used in the diamond drilling campaigns were used for the RC drilling, namely two in-house standards Std-4 and Std-6, as well as a commercial standard, AMIS185. A total of 285 standards were analysed, 95 for each standard. The results were assessed by MSA and values above three standard deviations outside of the certified mean were considered a failure. The standard results suggest a high degree of accuracy for the individual rare earth elements, with failure rates generally below 10%, with the exception of samarium for Std-4 having a 15% failure rate and thulium for Std-6 having a failure rate of 22%.

Control charts for dysprosium assays of the standards are presented in Figure 82 and a summary of the number of failures and the failure rate for each standard is presented in Table 30.

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Figure 82 Dysprosium standard control charts

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Table 30 Failure rate (outside ± 3 SD) for standard reference material assayed by Actlabs during the 2023 RC drilling campaign

CRM Name	Failure Rate	La	Ce	Pr	Nd	Sm	Eu	Gd	Tb	Dy	Ho	Er	Tm	Yb	Lu	Y
STD 4	Number of Samples	5	0	0	2	14	0	1	7	6	0	7	9	5	0	0
	Percentage	5%	0%	0%	2%	15%	0%	1%	7%	6%	0%	7%	9%	5%	0%	0%
STD 6	Number of Samples	9	1	7	1	0	0	0	2	2	6	0	21	9	0	0
	Percentage	9%	1%	7%	1%	0%	0%	0%	2%	2%	6%	0%	22%	9%	0%	0%
AMIS0185*	Number of Samples	0	0	0	0	0	1	-	-	0	0	-	1	-	-	0
	Percentage	0%	0%	0%	0%	0%	1%	-	-	0%	0%	-	1%	-	-	0%

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11.5.3.3. Pulp Duplicates

Pulp duplicates were inserted into the RC sampling stream at a rate of 1 in every 20 samples. The original and duplicates samples were compared graphically and statistically to determine the precision of the assay results. A summary of the repeatability of the pulp duplicates is presented in Table 31 for TREO, LREO and HREO as well as Dy₂O₃. The results show a high degree of analytical precision, with nearly 99% of the duplicate pairs reporting a HARD value of less than 10%.

Table 31 Summary of lab duplicate repeatability

Variable	Number of Samples	Original Mean	Duplicate Mean	Percentage Difference	HARD
					< 10%	< 20%
TREO %	285	0.08	0.08	3%	98.9	99.6
LREO %	285	0.05	0.05	4%	98.9	100
HREO %	285	0.03	0.03	2%	98.9	99.6
Dy₂O₃ ppm	285	31.9	32.7	2%	98.6	98.6

Scatter plots comparing the original with the duplicate assays are presented in Figure 83 for TREO and Figure 84 for Dy₂O₃.

Figure 83 Scatterplot of TREO assays in 285 pulp duplicates

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Figure 84 Scatterplot of Dy₂O₃ assays in 285 pulp duplicates

11.5.3.4. Field Duplicates

A total of 35 field duplicate samples were collected during the RC drilling campaign to assess the precision of the sampling process. Acceptable precision was demonstrated for TREO, LREO and HREO as well as Dy₂O₃, with over 90% of the duplicate pairs having a HARD value of <10% (Table 32).

Table 32 Summary of field duplicate repeatability

Variable	Number of Samples	Original Mean	Duplicate Mean	Percentage Difference	HARD
					< 10%	< 20%
TREO %	35	0.24	0.24	2%	97.1	97.1
LREO %	35	0.13	0.14	3%	94.3	100
HREO %	35	0.11	0.11	1%	94.3	97.1
Dy2O3 ppm	35	110.2	107.6	-2%	97.1	97.1

Scatterplots of the field duplicate pairs for TREO and Dy₂O₃ are shown in Figure 85 and Figure 86 respectively.

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Figure 85 Scatterplot of TREO assays for 35 field duplicates

Figure 86 Scatterplot of TREO assays for 35 field duplicates

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11.6. Adequacy of Sample Preparation, Security and Analytical Procedures

All aspects of core handling, marking, logging, cutting, bagging, labelling and sample submission to Actlabs' preparation facilities at Windhoek are covered by well-designed protocols to ensure that all routine activities are conducted with maximum consistency and followed industry standards.

Drillhole core handling and storage as well as core sampling and transport are conducted in a safe and secure manner. NMI followed an auditable chain of custody, which ensured high levels of security and integrity of the results.

The QP is of the opinion that the sampling and analytical procedures and the number of QAQC samples inserted into the sample stream are appropriate for the current level of the project, the type of the deposit and for the analytical techniques used. The blank sample, standard reference material and duplicate data show minimal contamination, acceptable accuracy and a high level of precision.

Field duplicates for the RC drilling indicate that the procedures are appropriate for the Lofdal mineralisation.

The analytical results from secondary laboratory confirm the primary laboratory results with acceptable limits.

It is the QP's opinion that the sample assay results are acceptable for use in a Mineral Resource Estimate.

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12. Data Verification

The Lofdal site was visited by the QP for the Mineral Resource (Jeremy Witley) from the 28th to the 30th of October 2020. During the site visit, the following verification work was completed:

The exploration processes were examined, and it was found that the work is being carried out according to the Lofdal procedures, which are appropriate for the purposes of evaluating the Mineral Resource.
The logging, sampling and assay records were examined for a selection of drillholes from both previous and 2020 drilling for Area 4 (nine pre-2020 holes and eight recent holes with ICP assay results) and Area 2B (13 recent holes, three with ICP assay results), and were verified against observations made on the cores. The logging was found to be of good quality and the higher grade REE mineralisation was observed to be associated with the lithology, alteration and structures as described in sections 7.4 of this report.
Ad-hoc hand-held XRF readings were taken on several cores that confirmed the presence of elevated Y. Although the results of this exercise are not definitive, it served to verify the magnitude of the assayed Y grades.
RadEye readings were taken on the core at the drilling rigs, at the field camp and core yard, as well as on outcrops of carbonatite dykes in the field. These readings confirmed the elevated gamma readings associated with the mineralised zones.
Cores were observed being taken from the drillholes during the drilling process for drillhole L4D0148 and L4D0151. The cores observed being removed from the holes exhibited albitisation and iron alteration in the footwall to the mineralisation at Area 4.
The drilling locations of completed drillholes in the 2020 and previous programs were observed in the field. Handheld GPS readings of the collar positions were taken for seven of the 2020 Area 4 drillholes and eight of the 2020 Area 4 drillholes. The handheld GPS coordinates were compared with the final DGPS surveys, and no material discrepancies (>5 m) were noted.
The general site and the carbonatite dyke outcrops were examined. The outcrops observed are generally aligned with the mapping performed by NMI and its predecessors.

The Lofdal site was again visited by the QP for the Mineral Resource (Jeremy Witley) on the 10th of November 2022. Three of the drillholes completed since the previous site visit, that were included in the 2020 Mineral Resource, were inspected as well as the bulk sampling pit from which the metallurgical samples were extracted.

Additional verification comprised:

Spot checks of the database against the original borehole logs.
Spots-checks of database against original assay certificates.
Examination of database used for mineral resource estimation for any errors.

The most recent personal site inspection was completed by the QP from the 21st to the 22nd of November 2023. The following work was undertaken:

The RC drilling at Area 4 was observed for drillhole L4R0218.
The RC drilling and sampling processes were demonstrated at the drilling rig and at the core yard.
The drilling locations of completed RC drillholes were observed in the field. Handheld GPS readings of the collar positions were taken for eleven of the 2023 RC drillholes. The handheld GPS coordinates were compared with the final DGPS surveys, and no material discrepancies (>5 m) were noted.

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The residual RC material storage was inspected at the Lofdal storage facility at Khorixas. In the opinion of the QP, the data verification processes completed, and observations made during the site inspections demonstrate that the data collected are adequate for the purpose of Mineral Resource estimation.

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13. Mineral Processing and Metallurgical Testing

13.1. Historical Testwork Background

Testwork has been ongoing for the Lofdal Rare Earths Project where early works, previously reported, informing metallurgical development for various studies and designs. These included:

Scoping metallurgical testwork on the Lofdal REE Deposit, Mintek, August 2013;
Phase 2 – Additional WHIMS testwork on the Lofdal High Grade REE Sample, Mintek, May 2013;
Phase 3 – Magnetic separation testwork on Lofdal High Grade and Low-Grade ores, Mintek, January 2013;
Rare earth characterisation testwork, Nagrom, December 2013;
Test Report of sorting rare earth elements oxides, Tomra Sorting Solutions, April 2013.
Ore Sorting at Rados and Tomra and IMS Engineering.
LDE laboratories gravity separation and magnetic separation testing.
Geolabs comminution characterisation.
SGS Flotation testing on sorted products with hydrometallurgical testing on flotation concentrates for production of mixed rare earth oxides.
SGS flotation variability testing and mineralogical studies
SGS flotation testing followed by hydrometallurgical testing on fresh low grade feed ore samples
Laboratory bench flotation testing
Locked cycle testing
Bulk flotation testing for downstream testing
5 tonne pilot flotation plant for bulk concentrate production
Slurry characterisation testing
Geochemical testing on flotation tailings
Bench top acid bake and water leach testing with impurity removal
Acid bake scale up with intermediate REE recovery testing
Mini acid bake pilot campaign and downstream testing to final REE precipitate.

13.2. Ongoing Testwork

A significant amount of metallurgical optimization test work continued, largely at SGS Lakefield, Canada, to provide additional design information to inform the optimization of the process plant. This followed on from earlier test work on which process design and previous project economic evaluations were based on.

Bulk ore sorting testwork was doe at Ondoto on bulk Lofdal ore samples using a full scale TOMRA XRT Ore Sorter.

13.2.1. Ore Sorting

Further ore sorting testwork has been carried out for Lofdal Prefeasibility (PFS) study to evaluate performance using technology advances. High rare earth losses posed a challenge for project economics from previous testing during preliminary economic evaluation (PEA).

Mineralization at Lofdal is amenable to XRT sorting by detection of higher density minerals which are co-genetic with xenotime. Results indicate that XRT sorting technology can provide

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significant upgrades to the run of mine ("ROM") by rejecting waste in form of albitite, gneisses, muscovite and chlorite schists.

One sorting tests were part of the company's value engineering during the PFS process "Lofdal 2B-4" for consideration only on upgrading the low-grade ore stream by XRT sorting prior to flotation, with high-grade ore supplied directly to flotation.

Initial tests with TOMRA's new AI based and deep learning application OBTAIN, yielded improved XRT sorting results as compared to previous test programs. This formed the basis for bulk sample test programs carried out by Gecko Namibia with the upgraded TOMRA sorter at the Ondoto LREE Mine in northern Namibia. The pilot-scale XRT test program was conducted on about 300 tons of run of mine ore in July-August 2025. Sorting tests were conducted separately on bulk samples from the hanging wall, the main ore zone and the footwall zone as these three zones are characterized by different host lithologies (gneisses, pegmatites, amphibolites) and mineralization pattern.

A total of 200 different test runs were conducted on a TOMRA COM Tertiary XRT at Gecko Namibia's facilities. The test work was conducted as a combination of two different image processing methods, Dual Energy and Inclusion Detection. A special Multi Density Class Model was applied to distinguish between six different sensitivities. For the inclusion detection TOMRA's deep learning-based classification CONTAIN™ was tested to detect visual patterns and textures to recover finely disseminated mineralization within the low contrast Lofdal material.

The -10mm size fraction was screened out and excluded from the ore sorting tests and the remainder of the respective samples screened into two size classes, namely a finer and coarser stream for separate sorting evaluation. The results were for both grain sizes for the low-grade footwall material. The coarser fraction performed more sufficiently, achieving a better recovery to upgrade ratio for HREO and low reduced waste grades. These findings highlight the potential of prioritizing coarser grain size fractions in future plant design layout to maximize recovery and reduce waste grades. Increasing the bottom size enables higher throughputs without increasing mass pull to product.

The test results were steadily improved through 27 test settings by systematically adapting the multivariate test principles, parameters and algorithms based on the results as the program advanced. While the nature of mineralization with fine veins of xenotime is not the ideal type of material for XRT sorting, the test results exceeded the targeted upgrade and recoveries. The overall test results on low-grade (0.10-0.17% TREO) footwall and hanging wall ore yielded REE upgrade factors of 2.3 to 2.7 and REE recoveries of 60% to 70%.

13.2.2. Physical Processing

The target of the flotation optimization program was to simplify the flowsheet developed in previous test work, with similar or better flotation performance and to establish trends in the ore variability with particular focus on the variability flotation test work and optimisation program.

For the previous pilot program, the flotation pilot plant was conducted for a total of 105 hours of operation spread out over various separate campaigns. The overall pilot plant setup included a grinding circuit (consisting of one rod mill and one ball mill), a size classification circuit (screen with aperture size at 74 µm), and a flotation circuit (three roughers and two cleaners). The final pilot plant flowsheet is presented in Figure 87. The flotation feed was targeted at a P80 of 38 µm, and the rougher conditioning pulp density was 28%. The pulp temperature of rougher circuit was at 30°C through the whole pilot plant testing. The pulp temperature of the cleaner circuit was at 50°C through most of the operation. The collector Florrea 3900Z was eventually replaced with Florrea 3900.

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Figure 87 Pilot Plant Flowsheet

The pilot plant testing results are summarized in Table 33. The total rare earth oxide (TREO) recovery in the 2nd cleaner concentrate in PP-07A to PP-08C ranged from 48.0% to 62.6%, with a grade ranging from 2.21% to 3.27% TREO, and a mass pull ranging from 2.5% to 5.0%. 93 kg of REE flotation concentrate with an average grade of 0.88%Y (~2.65%TREO) was used for hydrometallurgical test work.

The pilot plant test work not only confirmed the flowsheet in a continuous operation but also demonstrated similar metallurgical performance at simpler and thrifted conditions for simpler and lower cost process operation. These included lower density conditioning (simpler in-circuit slurry density management), new collector Florrea 3900Z at a much lower collector dosage (~1000 g/t less than Florrea 3900) and a lower pulp temperature at 30°C (versus 50°C in previous design) but required a higher dosage of collector Florrea 3000 compared to the lab testing.

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Table 33 Pilot Plant Testing Results Summary

PP No.	Weight %	Measured Assay %					% Distribution
		Y	TREO	CaO	Fe_{2}O_{3}	SiO_{2}	Y	TREO	CaO	Fe_{2}O_{3}	SiO_{2}
Benchmark*	3.2	0.95	-	4.52	50.9	17.0	59.5	-	1.8	24.6	1.1
PP-02	1.4	0.72	2.90	4.97	43.1	21.0	19.5	24.7	0.8	0.8	9.6
PP-03	0.6	1.42	5.62	4.03	50.1	11.7	20.9	19.0	0.3	0.3	4.5
PP-04	4.2	0.68	-	7.86	34.9	26.9	58.1	-	3.9	23.2	2.1
PP-06	2.5	0.90	-	6.96	37.0	24.6	43.2	-	2.1	15.1	1.2
PP-07A	2.5	1.11	3.27	7.06	42.6	18.3	55.6	48.9	2.1	17.3	0.9
PP-07B	5.0	0.75	2.21	7.92	34.5	26.6	69.0	62.2	4.9	27.6	2.5
PP-08A	2.6	1.07	3.14	5.51	43.0	20.9	52.1	48.0	1.7	18.3	1.0
PP-08B	4.7	0.74	2.32	6.22	40.3	24.3	68.3	62.6	3.5	30.8	2.1
PP-08C	4.1	0.76	2.29	6.59	36.8	27.3	62.2	55.9	3.3	24.7	2.1
Average of PP-07A to PP-08C	3.8	0.89	2.65	6.66	39.4	23.5	61.4	55.5	3.1	23.8	1.7

Benchmark* is an average of lab testing PPF1 to PPF6 and PPF10

For the variability program, nine different variability samples were initially evaluated, plus a second batch of samples for flotation variability testing. following mineralogical characterisation showing that

The first batch of samples included nine variability samples, Var 1 to Var 9, in which the total rare earth oxide (TREO) contents ranged from 0.16% TREO to 0.64% TREO. The content of yttrium, as the proxy of TREO content, ranged widely from 541 g/t Y₂O₃ to 1,308 g/t Y₂O₃. In the second batch of samples, the TREO grades in Var 10 to Var 16 were lower than samples in the first batch, ranging from 0.15% TREO to 0.41% TREO. Yttrium was the most abundant rare earth element in the second batch of samples, ranging widely from 521 g/t Y₂O₃ to 2,641 g/t Y₂O₃. Two mini composites that were generated later in the program, Comp 2B-1 and Comp 4-1, contained 0.22% TREO and 0.27% TREO, respectively, which was similar to the TREO grade in run-of-mine (ROM) sample, at 0.22% TREO. Yttrium was also the primary rare earth element, at 1,040 g/t Y₂O₃, and 1054 g/t Y₂O₃ in Comp 2B-1 and Comp 4-1, respectively.

Sixteen variability samples Var 1 to Var 16 were submitted for X-Ray Diffraction (XRD) analysis, and TESCAN Integrated Mineral Analyzer (TIMA-X) analysis at a grind size P100 of 53 µm. The primary rare earth mineral was xenotime, which ranged from 0.10% (Var 1) to 0.71% (Var 11). The concentrations of other rare earth minerals (mainly synchysite and monazite) also varied widely in the sixteen variability samples, from 0.03% (Var 10) to 0.76% (Var 6), and zircon ranged from 0.01% (Var 2) to 0.36% (Var 12). Generally, the first nine variability samples (Var 1 to Var 9) contained less xenotime as compared to the seven variability samples in the second batch (Var 10 to Var 16).

The main gangue minerals in the sixteen variability samples consisted of quartz/ feldspars (38.8% - 77.5%), calcite/dolomite (6.4% - 45.5%), and biotite/chlorite/muscovite (5.41% - 19.5%). The main iron-bearing minerals were amphibole/epidote (0.67% - 5.13%), and Fe-oxides (0.5% - 12.6%). The contents of gangue minerals were found to be very different across the sixteen variability samples.

Individual flotation testing of the variability was carried out, followed by testing on two composite samples. A composite was made up from each of the areas with Area 2B, made up with slightly higher HREO portion as well as higher MgO and CaO and Area 4 samples which closely resembled the ROM composite head grade.

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For the first batch of samples (Var 1 to Var 9), the yttrium recovery to the final concentrate ranged from 27.2% (test VF-13/ Var 5) to 69.2% (test VF-7/ Var 4) at a grade of 0.28% Y₂O₃ (~0.79% TREO) to 0.55% Y₂O₃ (~1.71% TREO). The upgrading was best, at ~28 times, in test VF-3 on Var 9 sample and the lowest upgrading of only ~2 times was in test VF-8 on Var 6 sample. For the second-batch samples (Var 10 to Var 16), the yttrium recoveries to the 2nd cleaner concentrate ranged from 39.2% (test VF-39/ Var 14) to 66.2% (test VF-31/ Var 13) at a grade of 0.76% Y₂O₃ (1.54% TREO) to 0.94% Y₂O₃ (2.67% TREO). Test VF-30 on Var 11 sample achieved the highest upgrading ratio of ~21 times, and the 2nd cleaner concentrate in this test graded 6.52% Y₂O₃ (~10.05% TREO) at an yttrium recovery of 49.4% and 2.3% mass pull.

Two mini composites, Comp 2B-1 and Comp 4-1, were generated later in the testwork program to represent the samples from Area 2B and Area 4 of the deposit. Further flotation testing on these two mini composites with the best flotation results presented in Table 2. The 2nd cleaner concentrate produced in test VF-54 from Comp 2B-1 sample graded 0.77% Y₂O₃ (~1.63% TREO) at an yttrium recovery of 72.3%. About 73% of the yttrium was recovered to the 2nd cleaner concentrate in test VF-57 from Comp 4-1 sample at a grade of 0.81% Y₂O₃ (~2.11% TREO). The upgrade ratios achieved in test VF-54 of Comp 2B-1 and test VF-57 of Comp 4 were 7.2 and 8.2 times, respectively. It is noted that test VF-54 was completed on a finer primary grind size of minus 38 μm, whereas the primary grind size in test VF-57 was minus 53 μm, but with a regrind on the rougher concentrates to P80 of 34 μm.

VF-49 on Area 4 composite still had a high loss in the cleaners, but approaching the ROM baseline. Further refinements were done with adjustments to increase Calgon with aim to reduce Ca content and lower the sodium silicate were tested, with adjustments in the cleaners on Area 4 composite. Tests VF-50 and VF-51 respectively where the Ca content was largely reduced with the increase in Calgon, although with yttrium losses.

Further tests VF-52 on composite Area 2B (smaller flotation test cell) and VF-53 on composite Area 4 with adjustments on collector dosages. VF-53 showed similar recovery, but higher grade as compared to VF-49. The variability samples tended to have lower grade concentrates. Refer to Figure 88. Yttrium losses in the cleaner circuits tended to be where the main losses occurred.

The effect of reagent dosages and different grind sizes were evaluated in the sixteen variability samples as well as the two mini composites. The key findings were:

The dosages of Calgon, Florrea 3900 and Florrea 3000 in the rougher flotation stages in tests with Var 1 to Var 9 averaged 50 g/t, 1,200 g/t, 40 g/t, respectively; and those used in the cleaner stages were 10 g/t, 200 g/t, and 10 g/t, respectively.
The dosages of sodium silicate, Calgon, Florrea 3900, and Florrea 3000 used in the roughers in flotation tests with Var 10 to Var 16 were 250 g/t, 25-50 g/t, 1400 g/t, 80-120 g/t, respectively. And the dosages in the cleaners were 75 g/t, 5-10 g/t, 200-250 g/t, and 10-20 g/t, respectively.
For Comp 2B-1, the optimum dosages of sodium silicate, Calgon, Florrea 3900 and Florrea 3000 used in the roughers of test VF-54 were 150 g/t, 125 g/t, 1400 g/t and 100 g/t, respectively. For Comp 4-1, the optimum dosages of sodium silicate, Calgon, Florrea 3900 and Florrea 3000 in the roughers of VF-57 were 150 g/t, 60 g/t, 1400 g/t and 80 g/t, respectively.
The dosages of sodium silicate, Calgon, and collectors were critical to the flotation performance.
The flotation tests performed better at the finer grind size, minus 38 μm versus minus 53 μm. Alternatively, a primary grind at minus 53 μm followed by a regrind on the rougher concentrates to a P80 of 26-34 μm produced similar results as a primary grind of minus 38 μm.

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Table 34 Best Flotation Results from Variability Testing

Test #	Sample	Temp, °C	Products	Wt	Assay, %					Distribution, %
				%	Y2O3	TREO, est	SiO2	Fe2O3	CaO	Y2O3	TREO	SiO2	Fe2O3	CaO
CP-108	ROM	35	2nd CI Con NonMags	3.6	1.80	4.13	35.5	22.7	7.1	69.1	69.1	2.6	1.6	19.3
			Head (Cal.)	100.0	0.09	0.21	51.1	6.5	9.6	100	100	100	100	100
VF-4	Var 1	50	2nd CI Con NonMags	4.9	0.36	1.97	21.6	21.2	17.0	43.7	43.7	1.9	18.5	12.5
			Head (Cal.)	100.0	0.04	0.22	55.6	5.6	6.7	100	100	100	100	100
VF-5	Var 2	50	2nd CI Con NonMags	10.3	0.33	1.33	22.0	8.2	30.0	57.3	57.3	4.8	17.1	20.1
			Head (Cal.)	100.0	0.06	0.24	47.4	4.9	15.4	100	100	100	100	100
VF-18	Var 3	50	2nd CI Con NonMags	5.6	0.61	2.62	29.0	20.7	9.1	61.5	61.5	3.2	18.9	5.3
			Head (Cal.)	100.0	0.06	0.24	49.9	6.1	9.5	100	100	100	100	100
VF-7	Var 4	50	2nd CI Con NonMags	7.1	0.55	1.71	35.5	14.3	16.4	69.2	69.2	4.4	15.5	20.7
			Head (Cal.)	100.0	0.06	0.18	57.6	6.5	5.6	100	100	100	100	100
VF-13	Var 5	50	2nd CI Con NonMags	5.9	0.28	0.79	25.8	18.6	15.8	27.2	27.2	4.2	14.1	5.7
			Head (Cal.)	100.0	0.06	0.17	36.0	7.8	16.4	100	100	100	100	100
VF-8	Var 6	50	2nd CI Con NonMags	22.6	0.22	1.06	13.7	21.8	20.4	41.0	41.0	10.1	26.0	35.0
			Head (Cal.)	100.0	0.12	0.59	30.7	19.0	13.2	100	100	100	100	100
VF-9	Var 7	50	2nd CI Con NonMags	5.7	0.55	3.07	33.6	30.9	7.5	66.7	66.7	3.4	18.0	8.7
			Head (Cal.)	100.0	0.05	0.26	56.0	9.9	4.9	100	100	100	100	100
VF-15	Var 8	50	2nd CI Con NonMags	6.6	0.66	1.56	28.2	16.8	11.6	46.3	46.3	4.5	14.5	6.8
			Head (Cal.)	100.0	0.09	0.22	41.9	7.7	11.3	100	100	100	100	100
VF-3	Var 9	50	2nd CI Con NonMags	2.0	1.73	4.41	23.3	21.0	10.5	55.9	55.9	1.5	6.2	1.2
			Head (Cal.)	100.0	0.06	0.16	30.9	6.7	17.1	100	100	100	100	100
VF-23	Var 10	50	2nd CI Con 1-2	5.2	0.99	1.68	18.1	13.8	20.9	45.8	45.8	2.5	10.2	7.4
			Head (Cal.)	100.0	0.11	0.19	37.5	7.0	14.7	100	100	100	100	100
VF-30	Var 11	50	2nd CI Con 1-2	2.3	6.52	10.05	28.5	26.3	4.3	49.4	49.4	1.0	16.3	1.6
			Head (Cal.)	100.0	0.31	0.47	63.9	3.7	6.4	100	100	100	100	100
VF-25	Var 12	50	2nd CI Con 1-2	8.4	0.85	1.67	26.0	12.5	21.1	43.7	43.7	4.3	21.5	17.6
			Head (Cal.)	100.0	0.16	0.32	50.9	4.9	10.1	100	100	100	100	100
VF-31	Var 13	50	2nd CI Con 1-2	4.0	0.94	2.67	24.7	38.4	4.1	66.2	66.2	1.8	22.4	2.2
			Head (Cal.)	100.0	0.06	0.16	55.4	6.8	7.4	100	100	100	100	100
VF-39	Var 14	50	2nd CI Con 1-2	5.5	0.76	1.54	49.1	14.9	2.1	39.2	39.2	4.9	11.5	1.7
			Head (Cal.)	100.0	0.11	0.21	54.2	7.0	6.8	100	100	100	100	100
VF-28	Var 15	50	2nd CI Con 1-2	5.8	1.59	2.48	31.9	18.4	9.4	54.6	54.6	3.3	20.5	8.0
			Head (Cal.)	100.0	0.17	0.26	56.3	5.2	6.9	100	100	100	100	100
VF-40	Var 16	50	2nd CI Con 1-2	3.1	1.78	3.61	32.5	19.1	5.9	43.9	43.9	2.7	12.1	1.0
			Head (Cal.)	100.0	0.12	0.25	36.8	4.8	18.3	100	100	100	100	100
VF-54	Comp 2B-1	50	2nd CI Con 1-2	9.9	0.77	1.63	32.9	26.8	9.6	72.3	72.3	8.6	25.8	6.5
			Head (Cal.)	100.0	0.11	0.22	38.0	10.3	14.8	100	100	100	100	100
VF-57	Comp 4-1	50	2nd CI Con 1-2	8.9	0.81	2.11	25.9	27.5	12.5	73.0	73.0	4.5	36.5	12.1
			Head (Cal.)	100.0	0.10	0.26	50.6	6.7	9.1	100	100	100	100	100

Figure 88 Flotation Variability Testing Additional Results

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Mineralogy studies were initiated on the cleaner tails to investigate possible explanations. Xenotime liberation was 72% and 62% (Figure 89) for Area 2B and Area 4 samples respectively.

Figure 89 Mineralogy - Liberation of Variability Flotation Cleaner Tailings

This prompted further flotation testing with finer grind and concentrate regrind, Test VF-54 to VF-57 taking into account the consideration that the cleaner tails were what floated in rougher circuit. These ultra fines with several microns tend to get lost in the cleaners. Objective is to deal with the liberation and also avoid overgrinding and excessive fines generation. Also, as the xenotime is so fine grained and often associated with other gangues, they become so sensitive to the depressants, so when depressants target gangues, they unavoidably cause the xenotime loss. Resulting in a balance for selectivity testing.

The finer grinds on both composites seem to be boosting the recovery and bring the reference lines closer to the ROM material. This demonstrates that a fine grind of regrind is critical to improved results, especially for Area 4 composite. Refer Figure 90.

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Figure 90 Flotation Variability Testing Finer Grinds

Table 35 Flotation Variability - Summary of Finer Grind Flotation

Comp 2B-1					Final Conc
Test	Condition	P100, um	P80, um	Regrind	Wt, %	Grade, Y2O3%	Recovery, %
VF-48	Baseline	53	38	N	8.8	0.81	67.6
VF-54	finer primary grind	38	30	N	9.9	0.77	72.3
VF-56	regrind	53	38	Y	8.2	0.89	67.7

P80 of VF-56 regrind = 26 um

Comp 4-1					Final Conc
Test	Condition	P100, um	P80, um	Regrind	Wt, %	Grade, Y2O3%	Recovery, %
VF-53	Baseline	53	38	N	10.4	0.60	60.6
VF-55	finer primary grind	38	34	N	14.2	0.56	77.1
VF-57	regrind	53	38	Y	8.9	0.81	73.0

P80 of VF-57 regrind = 34 um

Following the variability work, recommendations were made for further testwork:

Investigate alternative depressants to depress carbonate and iron oxide gangue minerals more effectively.
Further investigate the flotation performance at a primary grind size P100 = 38 µm or with a regrind.
Investigate new flotation technologies for recovering minerals with fine particle size.
Re-evaluate magnetic separation as a pre-concentration step to produce a more consistent flotation feed and reduce the mass reporting to the flotation circuit, thereby minimizing reagent consumption and comminution costs.

Additional flotation testing was done on remaining ROM Bulk sample in baselining with deionised water; testing diluted Florrea collectors, confirming finer grind size of -38 µm; testing Axis House collector (at different dosages) as an alternative to Florrea collectors.

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Figure 91 Further Flotation Results on ROM Bulk Sample

Further optimization testwork on the ROM Bulk-1 sample indicated that a finer grind size at minus 38 µm presented better flotation results compared to a grind size at minus 53 µm.
The emulsification of Florrea collector 3000 is expected to have an impact on its effectiveness, but results are not conclusive. Further confirmation is required.
The flotation results using alternative collector AM810 from Axis House were not as good as using Florrea 3000 even at a slightly higher dosage.

Flotation testing was carried out on sorted samples. This to confirm the developed flowsheet on the three ore sorted samples (B13 FW, B15 HW, and B4 Main).

The total rare earth oxide (TREO) grade of the three ore sorted samples were 0.28% TREO for B13 FW, 0.40% TREO for B15 HW, and 0.23% for B4 Main. The yttrium content (Y) ranged from 895 g/t Y₂O₃ (B13 FW) to 1,778 g/t Y₂O₃ (B15 HW). The impurities in these samples were mainly SiO₂ (29.1%-58.5%), CaO (6.21% - 19.0%), Fe₂O₃ (5.92% -12.6%), and MgO (0.51% - 5.73%).

The response of the three sorted samples to flotation was reasonably good with minor modification to the reagent dosages. Furthermore, two bulk flotation tests using 10 kg charges were performed on the combined sample (Comb SP-2025) made from the three sorted samples. Test CP-202 results were better than the small batch flotation tests performed on the individual sorted samples. The selected best flotation test results are summarized in Table 36 and depicted in Figure 92.

Table 36 Selected Best Flotation Test Results on Sorted Feed Samples

Test #	Sample	Products	Wt %	Assay, %					Distribution, %
				Y₂O₃	TREO	CaO	SiO₂	Fe₂O₃	Y₂O₃	TREO	CaO	SiO₂	Fe₂O₃
F3	ROM Bulk-1	2nd CI Con	2.2	1.59	4.32	5.5	14.9	49.2	56.0	56.0	1.5	0.6	17.8
F209	B13 FW	2nd CI Conc NonMags	3.2	1.85	5.71	3.6	28.7	34.6	69.4	69.4	1.8	1.6	18.1
F210	B15 HW	2nd CI Conc NonMags	5.5	1.91	4.25	10.6	12.6	43.5	63.2	63.2	3.1	2.4	18.7
F208	B4 Main	2nd CI Con	4.0	1.59	3.48	7.4	13.7	51.6	57.8	57.8	2.2	1.3	17.6
CP202	Comb SP-2025	2nd CI Con NonMags	5.9	1.66	3.89	13.2	17.6	34.9	77.9	63.5	6.1	2.4	20.5

Note the TREO grade is measured for test CP202 only, the rest are estimated based on Y grade

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Figure 92 Selected Best Flotation Results of Sorted Feed Samples

The following concluded from the flotation testing on sorted samples:

The TREO grade of the three ore sorted samples ranged from 0.23% TREO (B4 Main) to 0.40% TREO (B15 HW). The yttrium content (Y) ranged from 895 g/t Y₂O₃ (B13 FW) to 1,778 g/t Y₂O₃ (B15 HW). The sorting bypass fines, Lofdal SF-2 sample, contained lower TREO grade, at 0.18% TREO and 696 g/t Y₂O₃.
The developed flowsheet was successfully applied to the three ore sorted samples (B13 FW, B15 HW, and B4 Main). With limited 1-2 tests per sample, the flotation results achieved were reasonable and comparable with previous samples.
Using a combined feed from the three sorted sample, the bulk flotation test results of CP-202 in 10 kg charge were good, better than the flotation tests on the individual sorted sample in 2 kg charges. This might be related to the lower pulp density in the rougher stage that was used for the 10 kg flotation test.

Evaluation testing for pre-concentration potential for the sorted bypass fines sample (Lofdal SF-2) by magnetic separation was considered. The magnetic separation testing on the sorted bypass fine Lofdal SF-2 sample performed effectively, with the combined magnetic concentrate grading 0.13% Y₂O₃ (~0.33% TREO) at 90.5% yttrium recovery and a 51% mass yield, corresponding to an overall upgrading ratio of approximately two times.

Further beneficiation testing recommendations were made:

Further optimize the flotation testing on the three sorted samples.
Evaluate the effect of lower pulp density in both rougher and cleaners.
Evaluate the magnetic separation on the run of mine ore and compare its performance with sorting.
Perform flotation testing on the sorted sample combined with the upgraded sorting fines.

13.2.3. Hydrometallurgical Processing

The successive optimization test work comprised the detailed downstream hydrometallurgical testing program, which aimed to facilitate effective scale up of the Acid Bake and Water Leach (ABWL) process and generate sufficient leach liquor to conduct a thorough investigation into optimizing downstream REE recovery steps.

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A series of tests was conducted focusing on parameters such as acid bake temperature, sulphation time, and two-stage sulphation:

Effect of acid bake temperature. Figure 93 shows the relation between sulphation temperature and Dy (used as a gauge for HREE extraction) and Fe extraction. The data shows a decrease in Fe extraction from 83% to 21% when the acid bake temperature was increased from 600°C to 650°C. This is coupled with a drop in Dy extraction of 93% to 89%. The reduction of iron extraction is attributed to the thermal decomposition of ferric sulphate to hematite and is expected to lead to a significant reduction in downstream MgCO₃ consumption.
Effect of sulphation time. The results show that Dy extraction did not change significantly with test time, however, extending the test time from one (1) to three (3) hours was favourable in suppressing iron from 54% Fe extraction to 41%, then 39%, respectively. It appears that a test duration of 2-3 hours is sufficient to achieve optimum Dy extraction and Fe rejection.
Effect of acid dosage (single stage sulphation). The impact of acid dosage is shown in Figure 94 and shows increased extractions of both dysprosium and iron at 1500 kg/t H₂SO₄ when compared to 1250 kg/t.
Effect of two-stage sulphation. Two stage sulphation was investigated and appeared to provide beneficial results by taking advantage of the high metal sulphation at low temperature, and the selective decomposition of impurities at high temperature acid. Figure 95 shows that at similar levels of iron rejection, the dysprosium extraction is slightly elevated when sulphation is conducted over two stages.
Effect of acid dosage (two stage sulphation). The acid dosage during a two-stage acid bake process was also investigated. The results presented in Figure 96 show that increasing the acid dosage by 250 kg/t (versus the flotation concentrate) led to a 4-5% increase in dysprosium extraction. The increase in iron extraction was only 3% by increasing from 1000 kg/t to 1250 kg/t, however a significant increase of 18% was observed when 1250 kg/t was increased further to 1500 kg/t.

Figure 93 Acid bake performance at different temperatures

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Figure 94 Acid bake performance using different acid addition at 600°C

Figure 95 Comparison between single and two-stage static acid bake process

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Figure 96 Effect of acid dosage on two-stage acid bake process

Continuous acid bake test campaigns were conducted using a 6-inch diameter pilot scale rotary kiln. The angle of the kiln and rotational speed were adjusted to achieve a high retention time of around 100 minutes. An internal temperature of 300°C for the low temperature acid bake and/or 625/650°C for the high temperature sulphation step was maintained, with a target acid-concentrate mixture feed flowrate of 2 kg/h.

It is known that at higher temperatures (600-700°C), sulphated iron minerals may decompose into SO₂/SO₃ (and report as off-gas) and insoluble hematite. Under these conditions REE sulphates do not yet decompose and therefore the following benefits could be realized:

As excess sulphuric acid is decomposed, WL liquors do not contain much free acid. This could lead to a lower MgCO₃ demand in Impurity Removal (IR).
Due to low solubility of hematite under water leach conditions, iron dissolution into the water leach liquor is significantly reduced (possibly from 60% to 25-30%). This could lead to a lower MgCO₃ demand in Impurity Removal (IR).
Thorium follows a similar pattern as iron and therefore, co-extraction of thorium is expected to be reduced as well.

Summarized results of the continuous kiln acid bake campaigns are presented in Figure 97 confirming that single stage sulphation yielded slightly lower dysprosium extraction (85-87%). Two-stage kiln acid bake achieved a Dy extraction of 92-94%. However, Fe extraction was significantly higher in the kiln run (90% and 72%) compared to static acid bake (61% and 32%), at 600°C and 650°C, respectively. The exact reasons for incomplete decomposition of sulphated iron are unknown but could possibly be attributed to the formation of agglomerates. The decomposition of ferric sulphate into hematite and SO₂ could possibly be slower than on the surface of these agglomerates.

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Figure 97 Comparison between single and two-stage dynamic acid bake process

The phases 2 and 3 of the program focussed on the leach liquor treatment for simplification and optimisation of the initial flowsheet Figure 98.

Figure 98 Initial block flow diagram of hydrometallurgical flowsheet before the PFS test work

The optimized processing route implemented the following changes to the flowsheet:

Replacing the partial purification with two stages (primary and secondary) of impurity removal to increase overall impurity removal including the complete removal of thorium

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Replacing crude REE precipitation, releaching and REE oxalate precipitation with two stages of REE carbonate precipitation.
Replacing REE calcination with drying.

Additional test work led to the decision to perform the impurity removal step into two stages: (1) primary neutralization, where the majority of iron, aluminium and thorium was removed at negligible HREE co-precipitation, and (2) secondary neutralization to remove the residual levels of iron, aluminium and thorium. Any co-precipitated HREE in the secondary neutralization step were to be recycled back to the water leach operation to be redissolved.

In the primary neutralization (PN) step, the goal was to remove the majority of the iron (and thorium), at minimum HREE co-precipitation and to improve the filtration behaviour. Several PN tests were conducted comparing different precipitation temperature (50°C, 90°C), initial Fe content in the feed (15.8 g/L, 21.6 g/L), pH conditions, and feed stream (leach filtrate or leach slurry). Based on the test work results, the following observations and conclusions could be made:

Effect of temperature on filtration. Generally, tests conducted at 90°C led to faster (280-400 kg/m2h) filtration rates than tests conducted at 50°C (30-80 kg/m2h). It is believed that at higher temperature the formation of goethite is favoured over the formation of amorphous ferric hydroxide. Filtration tests should ultimately be repeated using pulp samples produced under continuous precipitation conditions, as different morphologies may be produced and seed/recycling strategies could also then be considered. This will be considered during future studies.
Effect of temperature on metal precipitation. The results (Figure 99) show that complete iron and thorium removal was achieved regardless of precipitation temperature; while aluminium and dysprosium (based on residue assay) seemed to precipitate to a larger extent at higher temperature than at lower temperature. It is important to minimize dysprosium precipitation as any co-precipitated HREE with the PN residue are lost. Further tests to confirm dysprosium losses at high precipitation temperature were conducted. The values are reflected by hollow markers in
Figure 99b (PN-05) and Figure 99d (PN-06), respectively. The results show that dysprosium losses can be minimized to 2.5% and 4.1% at 15.8 g/L and 21.6 g/L Fe, respectively.
Effect of Fe concentration and temperature on MgCO3 consumption. Lower initial Fe content (15.8 g/L) required about 425 kg/t (concentrate) of MgCO₃ regardless of precipitation temperature. For tests with higher initial Fe content (21.6 g/L), the results show that precipitation test at 90°C requires more MgCO₃ (about ~200 kg/t more) than the precipitation test at 50°C. It's unclear exactly what mechanism contributed to this, but possibly the formation of different iron compounds (with varying degree of sulphur contents) contributed to the different consumption rates.
Effect of neutralization of leach pulp (vs filtrate). Neutralization behaviour of WL slurry and WL filtrate led to a total MgCO₃ consumption of 249 kg/t and 479 kg/t, respectively. This can be explained by the higher initial pH of the slurry (pH 1.2) compared to pH of the filtrate (pH 0.8). It can also be observed that under the same pH conditions, metal in the WL filtrate precipitates more significantly (see Figure 100). In the pulp, the leach residue acts as seed, promoting precipitation of impurities (over precipitation of REE). In filtrate, this effect is not present and precipitation of REEs takes place to a small extent.

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Figure 99 Metal precipitation as a function of pH and temperature

Figure 100 Metal precipitation behaviour of WL slurry and WL filtrate

In the secondary neutralization (SN) step, the goal is to remove residual impurities such as iron, aluminium, and thorium, producing a filtrate that is clean of impurities. The SN residue is recycled back upstream, so any co-precipitated HREEs are not lost. The first test was conducted in a titration-style where pH was incrementally increased from 3 to 5 using 10% w/w MgCO3 as neutralizing agent; the precipitation curve is presented in Figure 101.

The results show that at pH 5 most of the impurities (98.3-99.9%) are removed without significant co-precipitation of Dy (3.8%). The final filtrate (see Table 37) is clean (below detection limit for thorium, aluminium, and iron, and 1.8 mg/L uranium) and after uranium IX would be appropriate as feed for direct rare earth carbonate precipitation.

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Figure 101 Secondary neutralisation of pH

Table 37 Summary of secondary neutralisation tests

Metal	Assay, mg/L		Metal precipitation, %
	Feed	Final Filtrate
Dy	73	71	3.8
Th	3.3	<0.03	97.1
U	2.6	1.8	11.4
Al	72	<0.8	98.3
Fe	237	<0.2	99.9

For the uranium ion exchange, different resins were tested to remove residual uranium levels from the SN filtrate. For each test, 250 mL of feed was contacted with 1 mL of resin for a duration of 24 hours. The results show that Lewatit TP260 MDS (99%) and Puromet MTA4601PF (97%) removes the most uranium with low co-extraction of Dy (see Table 38). It is expected that any co-extracted Dy will be crowded-off at higher uranium loading once the resin is nearer to capacity. Kinetic tests showed that within 4 hours, more than 50% of uranium has been loaded to the resins, and after 24 hours, uranium loading was about 2 g/L (on resin). Between the two resins, Puromet MTA4601PF loaded the uranium faster. Based on these results, Puromet MTA4601PF was selected for bulk processing of the SN filtrate through a set of lead-lag IX columns. The composition of the UIX barren is shown in Table 39.

Table 38 Summary of the uranium IX results

Test ID		UIX1-1	UIX1-2	UIX1-3	UIX1-4
Resin		Lewatit TP260 MDS	Purolite S-930	Puromet MTA6002PF	Puromet MTA4601PF
Filtrate Assay, mg/L
Metal	0 h	24 h	24 h	24 h	24 h
Dy	82.4	62.9	73.4	78.8	82.4
U	1.96	0.02	0.15	0.09	0.07
Extraction, %
Metal	0 h	24 h	24 h	24 h	24 h
Dy	-	26	13	8	2
U	-	99	93	96	97

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Table 39 Composition of the U IX barren solution

| La
mg/L
149 | Ce
mg/L
261 | Pr
mg/L
25.8 | Nd
mg/L
86.5 | Sm
mg/L
22.3 | Eu
mg/L
9.07 | Gd
mg/L
47.6 | Tb
mg/L
10.8 | Dy
mg/L
75.3 | Ho
mg/L
16.5 | Y
mg/L
498 | Er
mg/L
46.2 | Tm
mg/L
6.32 | Yb
mg/L
35.8 | Lu
mg/L
4.75 |
| --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- |
| Sc
mg/L
<0.07 | Th
mg/L
<0.03 | U
mg/L
<0.02 | Si
mg/L
15.6 | Al
mg/L
<0.8 | Fe
mg/L
<0.2 | Mg
mg/L
12600 | Ca
mg/L
798 | Na
mg/L
38 | K
mg/L
<4 | Ti
mg/L
<0.2 | P
mg/L
<5 | Mn
mg/L
300 | Cr
mg/L
<0.1 | V
mg/L
<0.2 |

Rare earth precipitation (RP) was conducted to optimize and confirm test conditions before treating the remainder of the U-free liquor through a pilot plant. Result of initial precipitation tests from an SN filtrate is shown in Figure 102.

Figure 102 Dysprosium precipitation and tenors as a function of pH

The pilot plant with the optimized design operated for a total of 96 hours and consisted of two circuits – a primary RP (PRP) and a secondary RP (SRP) circuit. The purpose of the split RP circuits was to produce a higher purity (i.e. low Mg/Ca/Mn) in the primary RP circuit at slightly reduced efficiency (80-90%) with residual REEs recovered in the secondary RP circuit producing a RP precipitate with elevated levels of Mg, to be recycled back upstream to the leach discharge.

Figure 103 Rare earth precipitation pilot plant flowsheet

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The results in Table 40 show that the MREC products from lower (85-89%) and higher (95%) precipitation target yielded a product containing around 53% TREE, with slightly higher concentrations in the higher precipitation target. The impurities (Mg, Ca, Mn) levels are about the same except for slightly higher thorium (1.8 g/t) in the lower precipitation target than <0.5 g/t Th in the higher precipitation target. Minor thorium carbonate complex solubility at elevated pH may have contributed to this effect.

Table 40 Composition of PRP and SRP products generated in Pilot Plant

Circuit	PRP	PRP	SRP
Dy Ppt.%	85-89	95	99.8
La %	6.15	6.19	4.34
Ce %	11.4	11.0	5.26
Pr %	1.15	1.08	0.44
Nd %	3.96	3.71	1.42
Sm %	1.05	0.98	0.33
Eu %	0.42	0.39	0.15
Gd %	2.05	2.07	1.03
Tb %	0.47	0.44	0.24
Dy %	3.22	3.29	1.76
Ho %	0.70	0.66	0.44
Y %	18.2	19.7	23.1
Er %	1.97	1.98	1.27
Tm %	0.27	0.25	0.17
Yb %	1.49	1.40	0.86
Lu %	0.20	0.19	0.12
LREE %	24	23	12
HREE %	29	30	29
TREE %	53	53	41
Circuit	PRP	PRP	SRP
---	---	---	---
Dy Ppt.%	85-89	95	99.8
Sc g/t	<40	<40	<40
Th g/t	1.8	<0.5	4.9
U g/t	<0.5	<0.5	1.1
Si %	0.02	0.01	0.20
Al %	0.02	0.02	0.07
Fe %	<0.01	<0.01	0.11
Mg %	0.44	0.45	0.88
Ca %	0.16	0.20	4.56
Na %	0.18	0.19	0.09
K %	<0.01	<0.01	<0.01
Ti %	<0.01	<0.01	<0.01
P %	<0.01	<0.01	<0.01
Mn %	0.12	0.15	1.04
Cr %	<0.01	<0.01	<0.01
V %	<0.01	<0.01	<0.01

Based on the SRP performance, an overall Dy precipitation of 99.8% was achieved. Based on liquor assays, there is minimal Mg co-precipitation in the SRP, though SRP solids do contain slightly elevated levels of Mg (0.88%). The final SRP filtrate still has some residual dysprosium (~0.17 mg/L) and yttrium (5.6 mg/L minimum). In general, the SRP circuit operated as designed and seemed able to recover residual REE units from PRP filtrate. It is noted that the precipitate from SRP will be recycled to leaching and any REE contained in SRP solids will be recovered.

With recycling of SN and SRP precipitates, test work showed that any REE contained in the SN and SRP solids can be recovered using the excess acidity in the water leach liquor. Separate SN and SRP redissolution tests as well as combined SN/SRP redissolution tests were conducted. No additional acid was added to these tests. The tests showed that the SRP solids dissolved completely, and REE recovery from SN precipitates was calculated to be >99.7%. These tests confirmed that any REE contained in SN and/or SRP solids can easily be recovered without additional acid demand.

This program showed that a simplified acid bake and liquor treatment flowsheet consisting of a high temperature acid bake, two stage (primary and secondary) impurity removal, followed by UIX and two stages (primary and secondary) of HREE carbonate precipitation is able to produce a high grade HREE carbonate. The flowsheet developed in this program, presented in Figure 104, has eliminated several unit operations from the original flowsheet. The removal of crude REE precipitation, re-leach and thorium solvent extraction forms a significant simplification and contribution to an overall reduced reagent demand.

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Figure 104 Simplified flowsheet developed during further downstream hydrometallurgical testing

The following conclusions summarize key findings of the optimization test work program:

Under optimum operating periods, continuous high (600°C) temperature sulphation in a pilot rotary kiln yielded high HREE dissolution (94% Tb/Dy).
Batch test work was used to show that two stages of impurity removal using magnesium carbonate was able to remove practically all (below analytical detection limits) thorium, scandium, iron, aluminium and some of the uranium at minimum losses of HREE (~2%).
Uranium was removed by ion exchange using a conventional strong base anion resin (Puromet MTA4601PF). Uranium levels were reduced to below detection limit (0.02 mg/L U) with negligible co-extraction of HREE.

The U IX barren liquor was used in a mini pilot plant where a HREE carbonate was produced.

The circuit consisted of two stages (primary and secondary) of precipitation using sodium carbonate. Overall recovery of HREE over two stages was almost quantitative and around 0.5 kg of HREE carbonate precipitate was produced at 53% TREE (3.25% Dy, 0.45% Tb, 19.0% Y, 1.12% Pr, 3.83% Nd) and typical impurity levels of <0.5 g/t U, <0.5 g/t Th as well as 0.44% Mg, 0.13% Mn and 0.18% Ca.

Solid-liquid separation tests were carried out on water leach residue (BW1-PR), impurity removal pulps (BIR3) as well as the primary and secondary rare earth precipitation pulps (test PP1 PRP3 and PP1 SRP3).

Dynamic thickening and rheology tests indicated that the BW1-PR Residue and BIR3 Pulp underflow samples both exhibited Bingham plastic behavior and were generally thixotropic. The BW1-PR Residue showed poor overflow clarity, with overflow TSS averaging around 300 mg/L even at high coagulant and flocculant dosages. Pressure filtration achieved lower residual cake moisture than vacuum filtration, while wash efficiencies exceeding 100% for calcium suggested partial solids dissolution during washing. Similarly, BIR3 Pulp required moderate reagent dosages to achieve good settling, but filtration performance was limited by cake cracking and

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wall separation, which hindered dewatering and reduced washing efficiency; calcium wash efficiencies again exceeded 100%, indicating solids dissolution. The PP1 PRP3 sample responded well to Magnafloc 338 in static settling tests and exhibited Bingham plastic and thixotropic behavior above 8.1% w/w solids. However, both vacuum and pressure filtration produced low throughput and high residual moisture, with calcium and dysprosium wash efficiencies above 100%, again suggesting solids dissolution. Dynamic thickening testing is recommended for this material when sufficient sample becomes available. In contrast, the PP1 SRP3 sample, with its low solids content, showed poor settling performance even at high flocculant dosages. Further testing was not pursued. Recommended to investigate alternative separation methods.

Overall, the test program showed that a simplified acid bake and liquor treatment flowsheet consisting of a high temperature acid bake, two stage (primary and secondary) impurity removal, followed by UIX and two stages (primary and secondary) of REE carbonate precipitation is able to produce a high grade HREE carbonate.

The flowsheet developed in this program, presented in Figure 104, has eliminated several unit operations from the original flowsheet (Figure 98). The removal of crude REE precipitation, releach and thorium solvent extraction forms a significant simplification and is contributing to an overall reduced reagent demand.

13.2.4. Further Testwork

Further testwork on flotation concentrates produced from the upgraded sorted low grade ore samples is still ongoing at time of writing.

13.3. Basis of Design

13.3.1. Recommendation for the Preferred Flowsheet

The testwork results for the physical separation and hydrometallurgical processing are discussed in detail in sections 13.2.1 to 13.2.3 respectively, including the reasoning behind selection of the various processing options and simplification of the flowsheet that was previously developed. Combining the physical separation and hydrometallurgical processing flowsheets results in an overall flowsheet which is illustrated in Figure 105. The applicable mass pulls and TREO recoveries that have been selected are summarised in Table 41.

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Figure 105 Recommended flowsheet

Table 41 Summary of selected recoveries
| | Mass pull | TREO recovery |
| --- | --- | --- |
| Low grade screening | 80% | |
| Ore sorting | 23% | 48.0% |
| Flotation | 3.4% | 64.4% |
| Hydrometallurgy circuit | | 92.2% |

13.3.2. Risks and Opportunities

Destruction of lixiviant and subsequent neutralization is costly in the hydrometallurgical flowsheets.

The sulphuric acid which is added to the sulphation roasting kiln is expected to be a major contributor to the operating cost of the hydrometallurgical processing circuit. The rate of addition that offered high REE recovery was chosen from the testwork results, however we believe that there is opportunity to optimise both the rate of acid addition and investigate means of recovering acid from the kiln of gas. In addition, the acid bake temperature and residence time could be further optimised to maximise REE recovery while minimising reagent consumptions.

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14. Mineral Resource Estimates

On behalf of NMI, MSA completed a Mineral Resource Estimate for the Area 4 and Area 2B deposits at the Lofdal Heavy Rare Earths project.

To the best of the QP's knowledge there are currently no title, legal, taxation, marketing, permitting, socio-economic or other relevant issues that may materially affect the Mineral Resource described in this Technical Report.

The Mineral Resources presented herein, with an effective date of 5 April 2024, represent an update to the previous Mineral Resource Estimate dated 12 May 2021. Drilling data derived from the 2023 RC drilling programme together with the diamond drilling data collected from 2010 to 2020 were used to update the previous estimates. It is the QP's opinion that the drilling data were collected in accordance with The Canadian Institute of Mining, Metallurgy and Petroleum (CIM) "Exploration Best Practices Guidelines", 2018.

The Mineral Resource was estimated using the 2019 CIM "Best Practice Guidelines for Estimation of Mineral Resources and Mineral Reserves" and classified in accordance with the "2014 CIM Definition Standards". It should be noted that Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

The Mineral Resource estimate was conducted using Datamine Studio RM and Leapfrog Geo software, together with Microsoft Excel, JMP and Datamine Supervisor for data analysis. The Mineral Resource estimation was carried out by Mr. Rui Goncalves under the supervision of Mr. Jeremy Witley (the Qualified Person).

14.1. Mineral Resource Estimation Database

The database provided by NMI to inform the Mineral Resource Estimate consists of:

Diamond drillhole (DD) data:
Collar surveys.
Downhole surveys.
Sampling and assay data.
Geology logs.
Specific gravity (SG) measurements.
Recovery and Rock Quality Designation (RQD) measurements.
Reverse circulation (RC) data consisting of:
Collar surveys.
Downhole surveys.
Sampling and assay data.
Geological logs.
Information from trench data.
Topographic surveys were provided as contours in GIS shapefile format.

The drillhole and trench data were provided in Microsoft Excel files that were extracted from a Microsoft Access database managed by NMI. The principal sources of information used for the estimate are exploration diamond drilling programmes conducted by NMI from 2010 to 2012, 2015 and 2020, and a RC drilling programme completed in 2023. The trench data were used to guide the position of mineralised veins near surface but were not included in the grade estimation due to concerns on the representativity of the sampling in the near surface environment.

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A total of 186 diamond drillholes and 44 RC drillholes were drilled and 28 trenches were dug at Area 4. One drillhole (NLOFDH4007) was excluded, due to the absence of downhole surveys. Thirteen drillholes located 800 m to the northeast of the main drilling area were too far from the main area and were excluded from the Mineral Resource. Additionally, seven holes drilled within the plane of the mineralisation at Area 4 were used in defining the mineralised wireframes but were excluded from the Mineral Resource estimate as these samples are not representative of the mineralised package. The dataset for Area 2B consists of 46 diamond drillholes, 12 RC drillholes and 25 trenches.

A summary of the drilling undertaken per project is presented in Table 42.

Table 42 Summary of Lofdal drilling campaigns

Project	Drilling Campaign	Drilling Type	Number of Drillholes	Drilled Metres (m)	Assayed Metres (m)
Area 4	2011	Diamond Drilling	47	4 190.66	1 953.13
	2012	Diamond Drilling	78	8 737.49	3 815.06
	2013	Diamond Drilling	5	709.27	669.83
	2020	Diamond Drilling	56	10 162.07	3 618.06
	2023	Reverse Circulation	44	9 043.00	4 203.00
	Total		230	32 842.49	14 259.08
Area 2B	2010	Diamond Drilling	13	1 547.12	462.09
	2011	Diamond Drilling	4	588.10	187.10
	2020	Diamond Drilling	29	4 400.48	1515.00
	2023	Reverse Circulation	12	1 772.00	635.00
	Total		58	8 307.7	2 799.19

The position of the drillhole collars for Area 4 are shown in Figure 106 (note that the 13 collars outside the Mineral Resource area are not shown). The collar positions for Area 2B are shown in Figure 107.

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Figure 106 Collar positions by campaign for Area 4

Figure 107 Collar positions by campaign for Area 2B

The cut-off date for inclusion of data into this estimate is 1 February 2024 at which time there was no outstanding information for Area 4 and Area 2B as the drilling was completed in 2023.

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14.2. Exploratory Analysis of the Raw Data

The dataset examined consisted of sampling and logging data from diamond drillholes and RC drillholes. The following attributes are of direct relevance to the estimate:

REE oxide grades in ppm: Lanthanum (La₂O₃), Cerium (C₂O₃), Praseodymium (Pr₂O₃), Neodymium (Nd₂O₃), Samarium (Sm₂O₃), Europium (Eu₂O₃), Gadolinium (Gd₂O₃), Terbium (Tb₂O₃), Dysprosium (Dy₂O₃), Holmium (Ho₂O₃), Erbium (Er₂O₃), Thulium (Tm₂O₃), Ytterbium (Yb₂O₃) and Lutetium (Lu₂O₃), as well as Yttrium (Y₂O₃).
Specific Gravity (SG) measurements derived from diamond drill cores.
Rock Quality Designation (RQD) measurements derived from diamond drill cores.

14.2.1. Validation of the Data

MSA undertook a high-level validation process which included the following checks:

Examining the sample assay, collar survey, down-hole survey and geology data to ensure that the data were complete for all the drillholes,
Examining the de-surveyed data in three dimensions to check for spatial errors,
Examination of the assay and density data to ascertain whether they were within expected ranges,
Checks for "FROM-TO" errors, to ensure that the sample data do not overlap one another or that there are no unexplained gaps in the sampling.
The data validation exercise revealed the following:
There are no unresolved errors relating to missing intervals and any overlaps in the drillhole logging data. Absent assays correspond to intervals where no samples were taken.
Examination of the drillhole data in three dimensions shows that the collars of the drillholes surveyed by DGPS plot in their expected positions relative to the topographic surface derived from the contour data.
Extreme assays were checked, and no errors were found.
Two methods were used to derive density measurements from the diamond drill cores. Density measurements on the drillhole data pre-dating the 2020 campaign made use of downhole geophysical probe surveys, while the 2020 campaign made use of the Archimedes principle on the drill cores. A statistical comparison for Area 2B between these two methods indicates that the downhole densities reported higher average values and are of a statistically different population (Figure 108). These differences were less pronounced for Area 4 but resulted in the downhole geophysical probe densities being excluded from the estimation process in favour of the Archimedes-principle measurements.
The average sample weight of each 1 m RC Sample was 28 kg, with sample weights ranging from 3 kg to 50 kg within the mineralised zone for Area 4 and 5 kg to 39 kg for Area 2B. Anomalousy heavy samples were recorded for Area 4 however these are located in the waste zone and have a minimal impact on the estimates.

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Figure 108 Cumulative frequency distribution comparison between pre-2020 downhole probe and 2020 Archimedes densities

14.2.2. Statistics of the Raw Sample Data

14.2.2.1. Sample Lengths

Diamond drillhole sample lengths vary from 0.1 m to 6.00 m in Area 4 and 0.19 m to 2.23 m in Area 2B with the dominant sample length being 1 m for both areas as illustrated in the histograms in Figure 109.

Figure 109 Histogram of DD sample lengths for Area 4 and Area 2B

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The majority of the RC samples were collected at 1 m intervals, with a small number of 0.5 m samples collected, typically at the end of a drillhole.

14.3. Bivariate Analysis

The relationships between individual rare earth oxides were studied using scatterplots to understand the existence of any correlation between variables which should be preserved in the mineral resource estimate. A strong linear relationship between the grades of certain REE exist, with some elements displaying this relationship with multiple elements. Linear relationships tend to be strongest for elements that are located adjacent to one another in the periodic table, with the exception of the Nd-Sm and Sm-Eu paired data, which marks the transition from light to heavy rare earth elements. As an example, Figure 110 shows the relationships of Tb2O3 with Dy2O3 and Ho2O3 for Area 4.

Figure 110 Scatter plot of sample $\mathrm{Tb_2O_3}$, $\mathrm{Dy_2O_3}$ and $\mathrm{Ho_2O_3}$ for Area 4

14.4. Core Recovery

14.4.1. Diamond Drillholes

The average core recovery is 93% for Area 4 and 95% for Area 2B. A broad depth-core recovery relationship exists showing increasing core recovery with increasing depth (Table 43). There is no discernible relationship between grade and core recovery.

Table 43 Diamond core recovery in percent per depth interval below surface

Area	Recovery in Percent per Depth Interval (m)
	0 – 5	5 – 10	10 – 20	20 – 30	30 – 40	40 – 50	Overall
Area 4	64.8	82.9	87.0	91.1	93.9	92.6	93.4
Area 2B	70.5	70.9	92.9	96.0	94.9	95.7	94.6

14.4.2. Reverse Circulation

The average recovered weight for the RC samples in Area 4 ranged from 1 kg to 130 kg. There are 10 samples in Area 4 with weights above 50 kg due to over-run drilled intervals, these occur in the waste zone and therefore have a minimum impact on the Mineral Resource. Recovered sample weights for Area 2B range from 1 kg to 65kg.

Summary statistics for the samples within the mineralised zone for each zone is presented in Table 44.

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Table 44 Summary statistics of RC sample weights
| Area | Number of Samples | Minimum (kg) | Maximum (kg) | Mean (kg) | CV |
| --- | --- | --- | --- | --- | --- |
| Area 4 | 3 473 | 3 | 50 | 27 | 0.25 |
| Area 2B | 370 | 5 | 39 | 30 | 0.19 |

Histograms for the recovered sample weights within the mineralised zone is presented in Figure 111 for Area 4.

Figure 111 Histogram of RC sample weights for Area 4 (left) and Area 2B (right)

No relationship between recovered sample weight and grade was observed.

14.5. Geological Modelling

Leapfrog Geo was used to generate three-dimensional volumes and surfaces representing the mineralised zones and weathering surfaces. The drilling data from the 2023 RC programme was used to update and adjust the existing geological model.

14.5.1. Topography

A topographic survey was provided by NMI which was conducted by UAS Flighttec Solutions (Pty) Ltd in 2020. This survey consists of topographic contours which were used to generate a three-dimensional surface in Leapfrog Geo. As no significant mining activities have taken place since this survey, the same topographical surface was used in the update of the 2023 Mineral Resource.

The surveyed drillhole collars correspond well with the resultant topographic surface. The trench data was draped onto the topographic surface, which was used to guide the modelling of the mineralised wireframes near surface.

14.5.2. Mineralised Zones

The modelling procedure followed the same methodology applied in the 2021 estimate. The visual continuity of dysprosium oxide $(\mathrm{Dy}_2\mathrm{O}_3)$ grades was examined along strike and down-dip to generate mineralised wireframes using a statistical threshold of $10~\mathrm{ppm}$ $\mathrm{Dy}_2\mathrm{O}_3$ for Area 4 and $12~\mathrm{ppm}$ $\mathrm{Dy}_2\mathrm{O}_3$ for Area 2B. The use of these thresholds resulted in generally continuous zones that form a suitable framework for block model grade estimation. The modelled zones (or domains) were individually coded into the drillhole data and volumes were generated using

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Leapfrog Geo. Where necessary, manual edits were incorporated to provide for geologically realistic shapes.

The 2023 RC data confirmed the previous interpretation, with only minor adjustments to the wireframes required. The changes largely impacted the minor mineralised zones that occur above the main zone which resulted in one additional domain modelled for Area 4, for a total of fifteen mineralised zones (Figure 112) and two for Area 2B, totaling nine zones (Figure 113).

Figure 112 Cross-section illustrating modelled mineralised zones for Area 4

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Figure 113 Cross-section illustrating modelled mineralised domains for Area 2B

14.5.3. Oxidation/Weathering Surface

Due to the lack of detailed visual weathering logging, the rock quality designation (RQD) values from the diamond drillholes were used as a proxy for weathering. The assumption being that lower RQD values will be associated with alteration due to weathering.

Log-probability plots were used to identify a RQD threshold value of between approximately 40 % to 45 % to represent the threshold on which to base a partially weathered surface. This threshold correlates well with areas near surface and highlighted zones of deeper weathering associated with structural features (Figure 114).

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Figure 114 Cross-section illustrating modelled RQD weathering surface for Area 4

14.6. Bias Test

As the quality of RC samples is more difficult to assess compared to diamond drill core and down-hole smearing can occur, a bias test between the DD and RC data was undertaken to determine the presence of any systematic differences in both grade and thickness between the two datasets. In this case, doing a global comparison is not effective as the DD data is concentrated in the high-grade zone of the deposit whereas the RC drilling targeted the lower grade peripheral areas. Therefore, a comparison was undertaken between the two closest drillhole pairs within the main mineralised zone (MZONE 1) for both deposits.

The results of the bias test show that for both deposits, grade and thickness differences show no bias to a particular drilling type (Table 45). Individual differences are most likely due to irregularity in the structurally hosted hydrothermal mineralisation type of deposit.

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Table 45 Bias test on Dy₂O₃ for Area 2B and Area 4

Distance between RC and DD pair (m)	Diamond Drillholes			RC Holes			Difference
	Hole ID	Dy₂O₃	Length	Hole ID	Dy₂O₃	Length	Dy₂O₃	Length
Area 4
36	L4D0155	47	39	L4R0207	34	56	-12	17
37	L4D0135	37	74	L4R0212	42	70	4	-4
51	L4D0152	69	47	L4R0218	46	50	-24	3
53	NLOFDH40 97	74	32	L4R0196	55	65	-19	33
56	NLOFDH40 96	84	59	L4R0213	74	75	-9	16
56	L4D0133	46	72	L4R0199	55	56	9	-16
58	L4D0133	46	72	L4R0198	78	57	32	-15
Area 2B
40	L2BD0051	24	21	L2BR006 3	31	17	7	-4
53	L2BD0051	24	21	L2BR006 0	35	37	11	16
56	L2BD0053	89	17	L2BR006 1	137	12	48	-5

14.7. Statistical Analysis of the Composite Data

Samples were composited to one metre lengths based on the dominant sample interval. Compositing was carried out inside the mineralised domain and statistics were analysed for the fifteen rare earth oxides. Log histograms of the composites for total rare earth oxides (TREO %), heavy rare earth oxides (HREO %), light rare earth oxides (LREO %) and dysprosium oxide (Dy2O3 ppm) are shown for MZONE 1 in Figure 115 for Area 4 and Figure 116 for Area 2B.

The following observations were made:

The distributions for the individual REO grades are positively skewed.
The CV for TREO is similar for both deposits.
Area 4 has a higher HREO grade than Area 2B - the average grade of Dy₂O₃ ppm in the main mineralised domain is 82 ppm for Area 4 and 67 ppm for Area 2B.

Both deposits show similar proportions of HREO and LREO in TREO (Table 46)

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Figure 115 Histograms of TREO, LREO, HREO and Dy₂O₃ ppm for MZONE 1 in Area 4

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Figure 116 Histograms of TREO, LREO, HREO and Dy₂O₃ ppm for MZONE 1 in Area 2B

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Table 46 Individual REO proportions for Area 4 and Area 2B

REO	Percentage of REO in TREO
	Area 4	Area 2B
La₂O₃	12.0%	10.8%
Ce₂O₃	21.6%	18.3%
Pr₂O₃	2.3%	2.0%
Nd₂O₃	8.4%	8.9%
Sm₂O₃	2.5%	3.9%
Total LREO	46.8%	44.0%
Eu₂O₃	0.9%	1.4%
Gd₂O₃	3.7%	4.9%
Tb₂O₃	0.8%	0.9%
Dy₂O₃	5.1%	5.6%
Ho₂O₃	1.1%	1.1%
Er₂O₃	3.1%	3.2%
Tm₂O₃	0.5%	0.5%
Yb₂O₃	2.8%	2.9%
Lu₂O₃	0.4%	0.4%
Y₂O₃	34.7%	35.1%
Total HREO	53.2%	56.0%

14.7.1. Cutting and Capping

An outlier analysis was completed on the composite data for the individual mineralised domains and capping was applied where applicable, cognisant of the bivariate relationship between rare earth oxides. The capping impacted between 1 and 15 samples in each domain for Area 4 and 1 and 19 samples for each domain Area 2B.

14.8. Geostatistical Analysis

14.8.1. Semi-Variograms

Experimental semi-variograms were calculated on the normal scores transformed composite data for total HREO and total LREO grades using Datamine Supervisor (previously Snowden Supervisor) software. Normalised semi-variograms were calculated so that the sum of the variance is equal to one.

The semi-variograms for Area 4 were updated with the 2023 RC data which resulted in a shorter range for the second structure. The new data did not result in any significant changes to the Area 2B semi-variograms. Due to the limited number of samples in the smaller domains, semi-variograms were modelled using only the MZONE 1 composites. The same model was applied to the other domains during estimation.

Variogram maps for Area 4 showed the presence of weak anisotropy with the longest direction along strike (065°). Double structured, spherical semi-variogram models were modelled for

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HREO while LREO grade continuity was modelled with three structures. Table 47 summarises the semi-variogram parameters for Area 4.

Data for Area 2B did not suggest the presence of anisotropy. Single structure, spherical models were fitted to the experimental points for both HREO and LREO. Table 48 summarises the semi-variogram parameters for Area 2B.

Semi-variogram models for HREO for Area 4 are presented in Figure 117.

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Table 47 Semi-variogram Parameters for Area 4

Attribute	Rotation Angle			Rotation Axis			Nugget Effect (C0)	Sill 1 (C1)	Range of First Structure (m)			Sill 2 (C2)	Range of Second Structure (m)			Sill 3 (C2)	Range of Third Structure (m)
	1	2	3	1	2	3			1	2	1		1	2	3		1	2	3
HREO	155	45	0	Z	X	Z	0.19	0.32	30	30	4	0.49	60	55	13	-	-	-	-
LREO	155	45	0	Z	X	Z	0.32	0.31	30	30	3	0.21	70	55	6	0.16	70	55	22

Table 48 Semi-variogram Parameters for Area 2B

Attribute	Rotation Angle			Rotation Axis			Nugget Effect (C0)	Sill 1 (C1)	Range of First Structure (m)			Sill 2 (C2)	Range of Second Structure (m)			Sill 3 (C2)	Range of Third Structure (m)
	1	2	3	1	2	3			1	2	1		1	2	3		1	2	3
HREO	140	50	0	Z	X	Z	0.21	0.79	60	55	5	-	-	-	-	-	-	-	-
LREO	140	50	0	Z	X	Z	0.31	0.69	60	55	5	-	-	-	-	-	-	-	-

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Figure 117 Semi-variogram models for HREO in MZONE 1 - Area 4

14.9. Block Modelling

Block models were generated for each project using 10 m by 10 m blocks in the X (easting) and Y (northing) direction and 5 m blocks in the Z (elevation) direction. The block model was not rotated.

Sub-celling was applied to optimally fill the modelled wireframes, resulting in minimum sub-cell of 2 m x 2 m x 1 m in X, Y and Z, respectively.

The common origins for the block models for Area 4 and Area 2B are shown Table 49.

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Table 49 Block model origins Area 4 and Area 2B
| Area | Easting (m) | Northing (m) | Elevation (m) |
| --- | --- | --- | --- |
| Area 4 | 469 500 | 7 752 800 | 500 |
| Area 2B | 466 900 | 7 754 300 | 600 |

14.9.1. Estimation Parameters

The search distance and rotation angles were based on the semi-variogram. Kriging Neighbourhood Analysis (KNA) was used to determine the minimum and maximum number of samples to be included in the search neighbourhood and the appropriate number of discretisation points to be used in a parent block. The KNA exercise looked at Kriging Efficiency as a metric of estimation quality and slope of regression was used to quantify the level of conditional bias when selecting the optimal parameters.

The search parameters are shown in Table 50 for Area 4 and Table 51 for Area 2B.

Table 50 Search Parameters for Area 4
| Attribute | Rotation Angles | | | Rotation Axis | | | Search Distance (m) | | | Number of Composites | |
| --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- |
| | 1 | 2 | 3 | 1 | 2 | 3 | 1 | 2 | 3 | Min | Max |
| HREO | 155 | 45 | 180 | Z | X | Z | 60 | 55 | 14 | 6 | 16 |
| LREO | 155 | 45 | 180 | Z | X | Z | 70 | 55 | 9 | 6 | 16 |

Table 51 Search Parameters for Area 2B
| Attribute | Rotation Angles | | | Rotation Axis | | | Search Distance (m) | | | Number of Composites | |
| --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- |
| | 1 | 2 | 3 | 1 | 2 | 3 | 1 | 2 | 3 | Min | Max |
| HREO | 140 | 50 | 0 | Z | X | Y | 60 | 55 | 5 | 6 | 18 |
| LREO | 140 | 50 | 0 | Z | X | Y | 60 | 55 | 5 | 6 | 18 |

Block grades were estimated in three passes, with the first pass using the search parameters shown in Table 50 and Table 51. The second search was expanded by a factor of 1.5 with a minimum of 6 and maximum of 16 samples included for Area 4, and a minimum of 6 and maximum of 18 for Area 2B. The third search made use of an expansion factor of 10, with a minimum of 5 and maximum of 16 samples included for Area 4 while Area 2B included a minimum of 5 and maximum of 18 samples in the search neighbourhood. Estimates using the third search parameter are of relatively low confidence with the parameters designed to estimate local average values.

Ordinary Kriging (OK) was used for the estimation of the rare earth oxides. The modelled semi-variogram and search parameters were applied to the individual rare earth oxides. Estimates were completed for each individual mineralised zone comprising of fifteen zones for Area 4 and nine for Area 2B.

Density was estimated independently for each zone using inverse distance weighting and applying the same search parameters as the HREO attribute. Where blocks where not interpolated with a density estimate, the average value of the zone was assigned.

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Dynamic anisotropy was used to align the search ellipsoids to account for local changes in the orientation of the mineralised zones along strike and dip. The dynamic search for each zone was orientated using trend surfaces created in Leapfrog Geo.

14.10. Validation of Estimates

The models were validated by:

Comparison of the global estimates against the average composite sample grades.
Swath plot validation.
Visual examination of the input data against the block model estimates.

The average grade of the block model for each individual zone was validated against the declustered composite grades (declustered to 100 mX by 100 mY by 20 mZ). Globally, the estimated block grades compare favourably to the input data, with relative differences of less than ten percent for the main mineralised zones. Larger percentage differences are noted for the smaller zones, which can be attributed to factors such the spatial arrangement and paucity of the data.

Swath plot validations in the X, Y and Z directions were used to locally validate the block estimates against the declustered sample composites. No material biases in the estimates of the individual elements were identified. Examples of a swath plot validation for MZONE 1 are shown for Dy₂O₃ in Figure 118.

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Figure 118 Swath Plot Validation for Dy₂O₃ – Area 4 MZONE 1

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The block model was examined visually to ensure that the drillhole grades were locally well represented by the model and it was found that the block grades validated reasonably well against the data. The model is less well locally representative of the data when extrapolating down dip, which was considered in the classification. Examples of this validation for Dy2O3ppm are illustrated for Area 4 (Figure 119) and Area 2 (Figure 120).

Figure 119 Area 4 block model cross-section coloured on $\mathrm{Dy}_2\mathrm{O}_3$

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Figure 120 Area 2B block model cross-section coloured on $\mathrm{Dy}_2\mathrm{O}_3$

14.11. Mineral Resource Classification

Classification of the Area 4 and Area 2B Mineral Resources was based on the degree of geological uncertainty, grade continuity and variability, frequency of the drilling data and the confidence parameter outputs from the kriging estimates. The main considerations in the classification are as follows:

All the data that inform the Mineral Resource have been collected by NMI, using acceptable principles and the assays passed the relevant QAQC tests.
The geological model is robust, and the grade shells exhibit good continuity with low variability within and between drilling sections.
Semi-variogram ranges for the attributes are more than the general drillhole spacing in most areas.
Given the aforementioned factors, the Mineral Resources have been classified using the following criteria:
The Mineral Resource was classified as Measured where the level of confidence in the estimates is high. This is underpinned by data on a drilling grid of $30\mathrm{m}$ spacing or less. The kriging efficiency is between $50\%$ and $80\%$ and the slope of regression is higher than 0.8 for the majority of the blocks in the model.
the Indicated Mineral Resource is underpinned by data on a drilling grid of approximately $50\mathrm{m}$ spacing. The kriging efficiency and the slope of regression are lower than for Measured and the drillhole spacing is too wide to interpolate grades to a high level of

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accuracy, despite drillhole spacing being within the modelled variogram range The Indicated areas are directly adjacent to the Measured areas.

the Inferred Mineral Resource was classified where the confidence for the estimates is low. In these areas the drillholes are sparse and local estimates cannot be reliably made. The Inferred area is directly adjacent to the Indicated areas and largely occur in the deeper portions and periphery of the Mineral Resource.

The classified block model for Area 4 is shown in Figure 121.

Figure 121 Mineral Resource classification for Area 4

Mineral Resources for Area 2B were classified as Indicated and Inferred in the same way as for Area 4. Areas that fall outside of this classification, where significant extrapolation of grades occurs beyond the data coverage, were not included in the Mineral Resource and were assigned a code of OOR (out of resource) in the model (Figure 122).

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Figure 122 Mineral Resource classification for Area 2B

The Mineral Resources could be affected by further infill drilling, which may result in increases or decreases in subsequent Mineral Resource estimates. Inferred Mineral Resources are high-risk estimates that may change significantly with additional data. It cannot be assumed that all or part of an Inferred Mineral Resource will necessarily be upgraded to an Indicated Mineral Resource due to continued exploration. The Mineral Resources may also be affected by subsequent assessments of mining, environmental, processing, permitting, taxation, socio-economic and other factors.

14.12. Mineral Resource Statement

The Mineral Resource estimate as of 5 April 2024 is presented in Table 52 for Area 4 and Table 53 for Area 2B. The Mineral Resource is stated at a cut-off grade of 0.10% total rare earth oxides (TREO) and reported from within a Whittle optimised pit-shell.

In the QP's opinion, the Mineral Resources reported herein at the selected cut-off grade have "reasonable prospects for eventual economic extraction", taking into consideration mining and processing assumptions.

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Table 52 Area 4, Measured, Indicated and Inferred Mineral Resource Estimates above 0.1% TREO cut-off grade – 5 April 2024

Category	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy₂O₃ ppm	TREO (kt)
Measured	6.6	0.21	0.14	0.07	130	13.7
Indicated	49.2	0.15	0.07	0.08	69	75.7
Measured & Indicated	55.8	0.16	0.08	0.08	76	89.4
Inferred	10.5	0.14	0.06	0.08	58	15.0

Notes:
All tabulated data have been rounded and as a result minor computational errors may occur.
Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
Quantities reported are the total quantities for the project regardless of ownership.
TREO = Total Rare Earth Oxides and includes Y₂O₃
HREO = Heavy Rare Earth Oxides and includes Y₂O₃
**LREO = Light Rare Earth Oxides
Mt = Million tonnes, kt = Thousand tonnes.

Table 53 Area 2B, Indicated and Inferred Mineral Resource Estimates above 0.1% TREO cut-off grade – 5 April 2024

Category	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy₂O₃ ppm	TREO (kt)
Indicated	2.7	0.16	0.09	0.07	97	4.4
Inferred	4.4	0.15	0.07	0.08	75	6.6

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The Mineral Resource for Area 4 is presented at a variety of cut-off grades as shown in Table 54 for the combined Measured and Indicated Resources and Table 55 for the Inferred Mineral Resource.

Table 54 Area 4, Measured and Indicated Mineral Resource grade-tonnage table – 5 April 2024

Cut-off TRE0%	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy2O3 ppm	TREO* (kt)
0.10	55.8	0.16	0.08	0.08	76	89.4
0.15	20.4	0.23	0.13	0.10	120	46.5
0.20	8.4	0.31	0.20	0.11	186	26.0
0.25	4.2	0.40	0.29	0.11	262	16.8
0.30	2.6	0.48	0.38	0.10	333	12.4

Notes:
All tabulated data have been rounded and as a result minor computational errors may occur.
Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
Quantities reported are the total quantities for the project regardless of ownership.
TREO = Total Rare Earth Oxides and includes Y₂O₃
HREO = Total Heavy Rare Earth Oxides and includes Y₂O₃
**LREO = Total Light Rare Earth Oxides
Mt = Million tonnes, kt = Thousand tonnes

Table 55 Area 4, Inferred Mineral Resources grade-tonnage table – 5 April 2024

Cut-off TRE0%	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy2O3 ppm	TREO* (kt)
0.10	10.5	0.14	0.06	0.08	58	15.0
0.15	3.4	0.18	0.08	0.11	76	6.3
0.20	0.7	0.24	0.12	0.12	118	1.7
0.25	0.2	0.30	0.20	0.09	193	0.6

Notes:
All tabulated data have been rounded and as a result minor computational errors may occur.
Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
TREO = Total Rare Earth Oxides and includes Y₂O₃
HREO = Total Heavy Rare Earth Oxides and includes Y₂O₃
**LREO = Total Light Rare Earth Oxides
Mt = Million tonnes, kt = Thousand tonnes.

The Mineral Resource for Area 2B is presented at a variety of cut-off grades in Table 56 for the Inferred Mineral Resource and Table 57 for the Indicated Mineral Resource.

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Table 56 Area 2B, Indicated Resources grade-tonnage table – 5 April 2024

Cut-off TRE0%	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy₂O₃ ppm	TREO* (kt)
0.10	2.7	0.16	0.09	0.07	97	4.4
0.15	1.3	0.21	0.11	0.10	117	2.7
0.20	0.6	0.25	0.12	0.13	133	1.5
0.25	0.3	0.29	0.14	0.15	150	0.8

Notes:
1. All tabulated data have been rounded and as a result minor computational errors may occur.
2. Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
3. Quantities reported are the total quantities for the project regardless of ownership.
4. TREO = Total Rare Earth Oxides and includes Y₂O₃
5. HREO = Total Heavy Rare Earth Oxides and includes Y₂O₃
6. **LREO = Total Light Rare Earth Oxides
7. Mt = Million tonnes, kt = Thousand tonnes.

Table 57 Area 2B, Inferred Resources grade-tonnage table – 5 April 2024

Cut-off TRE0%	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy₂O₃ ppm	TREO* (kt)
0.10	4.4	0.15	0.07	0.08	75	6.6
0.15	1.6	0.20	0.09	0.11	96	3.3
0.20	0.6	0.25	0.10	0.15	111	1.6
0.25	0.2	0.31	0.10	0.20	115	0.8

Notes:
1. All tabulated data have been rounded and as a result minor computational errors may occur.
2. Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
3. Quantities reported are the total quantities for the project regardless of ownership.
4. TREO = Total Rare Earth Oxides and includes Y₂O₃
5. HREO = Total Heavy Rare Earth Oxides and includes Y₂O₃
6. **LREO = Total Light Rare Earth Oxides
7. Mt = Million tonnes, kt = Thousand tonnes

The grades for the individual REE for each class are shown for Area 4 in Table 58 and in Table 59 for Area 2B.

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Table 58 Area 4, Individual REO Measured, Indicated and Inferred Mineral Resources above 0.1% TREO cut-off grade – 5 April 2024

Class	Tonnes Mt	TREO* %	La₂O₃ ppm	CeO₃ ppm	Pr₂O₃ ppm	Nd₂O₃ ppm	Sm₂O₃ ppm	Eu₂O₃ ppm	Gd₂O₃ ppm	Tb₂O₃ ppm	Dy₂O₃ ppm	Ho₂O₃ ppm	Er₂O₃ ppm	Tm₂O₃ ppm	Yb₂O₃ ppm	Lu₂O₃ ppm	Y₂O₃ ppm
Measured	6.6	0.21	173	313	34	124	42	18	81	19	130	28	83	12	76	11	935
Indicated	49.2	0.15	217	383	40	145	40	14	55	11	69	14	41	6	36	5	463
M&I	55.8	0.16	211	374	39	142	40	15	58	12	76	16	46	7	41	6	519
Inferred	10.5	0.14	217	389	42	150	40	13	49	9	58	12	34	5	30	4	369

Notes:
1. All tabulated data have been rounded and as a result minor computational errors may occur.
2. Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
3. Quantities reported are the total quantities for the project regardless of ownership.
4. *TREO = Total Rare Earth Oxides and includes Y₂O₃
5. Mt = Million tonnes, M&I is the summation of Measured and Indicated.

Table 59 Area 2B, Individual REO Measured, Indicated and Inferred Mineral Resources above 0.1% TREO grade – 5 April 2024

Class	Tonnes Mt	TREO* %	La₂O₃ ppm	CeO₃ ppm	Pr₂O₃ ppm	Nd₂O₃ ppm	Sm₂O₃ ppm	Eu₂O₃ ppm	Gd₂O₃ ppm	Tb₂O₃ ppm	Dy₂O₃ ppm	Ho₂O₃ ppm	Er₂O₃ ppm	Tm₂O₃ ppm	Yb₂O₃ ppm	Lu₂O₃ ppm	Y₂O₃ ppm
Indicated	2.7	0.16	187	303	32	126	51	20	73	15	97	19	55	8	51	7	596
Inferred	4.4	0.15	196	320	36	160	76	25	80	13	75	14	40	6	36	5	440

Notes:
1. All tabulated data have been rounded and as a result minor computational errors may occur.
2. Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
3. Quantities reported are the total quantities for the project regardless of ownership.
4. *TREO = Total Rare Earth Oxides and includes Y₂O₃
5. Mt = Million tonnes.

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14.13. Assessment of Reasonable Prospects for Eventual Economic Extraction (RPEEE)

In assessing "reasonable prospects for eventual economic extraction" (RPEEE) the Mineral Resource was reported from within a Whittle optimised pit shell using the following assumed parameters and a cut-off grade of 0.1% TREO.

Mining will be by open-pit methods:
55° slope angle in the partially weathered rock and 63° slope angle in the fresh rock
3% mining dilution
3% mining loss
10 m bench height
Ore production rate of 2.0 million tonnes per annum.
65% final metallurgical recovery of TREO
Costs were assumed as follows:
Mining cost for drill and blast: USD 2.65 / tonne mined.
Processing costs: USD 32 / tonne milled
G&A cost: USD 1.41 / tonne milled
Transport cost: USD 36.31 /tonne concentrate
Offshore treatment cost and shipment priced in discounted basket price.
NMI price USD 91.64 per Kg TREO+Y2O3 (based on the 2022 PEA) (Table 60).

Table 60 Distribution of TREO in Concentrate

REO	Pricing per REE	Distribution of individual REO in TREO % (concentrate values)
La_{2}O_{3}	$-	9.2%
CeO_{2}	$-	16.0%
Pr_{6}O_{11}	$201.00	1.7%
Nd_{2}O_{3}	$212.00	6.3%
Sm_{2}O_{3}	$5.00	2.2%
Eu_{2}O_{3}	$36.00	1.1%
Gd_{2}O_{3}	$109.00	4.3%
Tb_{4}O_{7}	$2 493.00	0.9%
Dy_{2}O_{3}	$587.00	6.2%
Ho_{2}O_{3}	$290.00	1.3%
Er_{2}O_{3}	$64.00	3.8%
Tm_{2}O_{3}	$20.00	3.5%
Yb_{2}O_{3}	$947.00	0.5%
Lu_{2}O_{3}	$17.00	42.4%
Y_{2}O_{3}	$500.00	0.6%

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A plan showing the extents of the block model in relation to the conceptual pit shell boundaries is shown in Figure 123 for Area 4 and Figure 124 for Area 2B and a section through the deepest part of the modelled pit shell is shown in Figure 125 for Area 4 and Figure 126 for Area 2B. The pit shell covers the majority of the Area 4 grade block model both aerially and at depth, however the narrower mineralisation at Area 2B resulted in the mineral resource being constrained at depth by the limits of the pit shell. The modelled pit shell areas lie entirely within ML 200 and the nearest boundary of the license is approximately $6\mathrm{km}$ to the north and $9\mathrm{km}$ to the east. The two pits are far enough away from each other to be operated as separate pits, although close enough so that ore will be transported to a central facility for processing. There is no infrastructure, such as major roads, power lines, water courses or settlements, within or within the immediate vicinity of the pit shell outline.

The reader is advised that the assessment of economic potential that is incorporated in the Mineral Resource is solely for the purpose of reporting Mineral Resources and does not represent an attempt to estimate Mineral Reserves.

Figure 123 Area 4 – plan showing block model relative to pit shell extents

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Figure 124 Area 4 section looking northeast showing block model relative to pit shell extents and topography

Figure 125 Area 2 – plan showing block model relative to pit shell extents

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Figure 126 Area 2 section looking northeast showing block model relative to pit shell extents and topography

14.14. Comparison with Previous Estimate

The Mineral Resource estimate detailed in this report represents the third Mineral Resource Estimate reported for Area 4 and the second Mineral Resource for Area 2B. A comparison between the previous estimate for Area 4, with an effective date 12 May 2021, and the current estimate is shown in Table 61 at a 0.10% TREO cut-off.

The total Mineral Resource for Area 4 increased from May 2021 to February 2024 as a result of the additional drilling which expanded the extent of the Mineral Resource along strike and down-dip from the previously defined area.

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Table 61 Area 4 – 12 May 2021 Mineral Resource Estimate compared with 5 April 2024 Mineral Resource Estimate

Classification	12 May 2021			5 April 2024
	Tonnes (Mt)	TREO* %	TREO (kt)	Tonnes (Mt)	TREO* %	TREO (kt)
Measured	5.93	0.21	12.71	6.57	0.21	13.7
Indicated	36.63	0.16	59.97	49.22	0.15	75.7
M&I	42.57	0.17	72.68	55.79	0.16	89.4
Inferred	6.09	0.17	10.12	10.52	0.14	15.0

Notes:
1. All tabulated data have been rounded and as a result minor computational errors may occur.
2. Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
3. *TREO = Total Rare Earth Oxides and includes Y₂O₃
4. Mt = Million tonnes, kt = Thousand tonnes, M&I is the summation of Measured and Indicated.

The Indicated Mineral Resource for Area 2B increased from May 2021 to February 2024 due to changes to the criteria used to determine optimised pit shell. The Inferred Mineral Resources increased due to additional drilling which expanded the Mineral Resource northeast along strike and down-dip.

Table 62 Area 2B – 12 May 2021 Mineral Resource Estimate compared with 5 April 2024 Mineral Resource Estimate

Classification	12 May 2021			5 April 2024
	Tonnes (Mt)	TREO* %	TREO (kt)	Tonnes (Mt)	TREO* %	TREO (kt)
Indicated	2.20	0.19	4.3	2.65	0.16	4.4
Inferred	2.58	0.19	4.8	4.37	0.15	6.6

Notes:
1. All tabulated data have been rounded and as a result minor computational errors may occur.
2. Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
3. *TREO = Total Rare Earth Oxides and includes Y₂O₃
4. Mt = Million tonnes, kt = Thousand tonnes, M&I is the summation of Measured and Indicated.

The reader is advised that the 12 May 2021 Mineral Resource estimates have been superseded by that of 5 April 2024 and is presented purely for comparative purposes.

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15. Mineral Reserve Estimates

15.1. Overview

During 2022 a Preliminary Economic Assessment (PEA) was completed for the Lofdal Heavy Rare Earths Project “2B-4” (Lofdal), the purpose of this study was to evaluate the Lofdal Mineral Resource, considering all techno-economic factors, and to determine if the Project was viable and if it was appropriate to progress towards a Prefeasibility Study (PFS) and estimate mineral reserves.

The PFS Project evaluation was undertaken based on the requirements of 2019 CIM “Best Practice Guidelines for Estimation of Mineral Resources and Mineral Reserves” and classified in accordance with the 2014 CIM “Definition Standards”. All Mineral Reserves estimated were based on this code.

The project evaluation is based processing 1,1 Mtpa High-Grade (≥ 0,16% TREO) ROM and 2,01 Mtpa Low-Grade ROM (0,10% ≥ TREO < 0,16%), producing between 1 850 and 2 700 tonnes per annum (tpa) of total rare earths oxide (TREO) in concentrate for sale from an open pit mining operation using comminution (crushing, milling), XRT-sorting, flotation and hydrometallurgical process. All required modifying factors and supporting infrastructure required to support the mining- and processing plan generated has been allowed for and are discussed in the following sections.

15.2. Project Tenure

The Lofdal property was initially held under Exclusive Prospecting License (EPL) 3400, which was granted in 2005. In November 2023, EPL 3400 was relinquished and succeeded by Mining Licence 200 (ML 200). Figure 127 illustrates the location of ML 200 within the boundary of EPL 3400 at the time of its relinquishment, as well as the current boundaries, roads, and the locations of the Hoppe Mineral Claims and Mining License Applications.

Figure 127 Location of ML 200 and EPL3400 at relinquishment

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15.3. Mineral Resources Considered for Mining

Mineral Reserve estimates for Area 2B and Area 4 are based on geological models and Mineral Resource Estimates (MRE) detailed earlier (Table 14 13), specifically from block models "a2bmod_20240215.dm." and "a4mod_20240215.dm". Only Measured and Indicated (M&I) resources were considered for the Mineral Reserve estimate. While Inferred Mineral Resources may be included in open pit designs and mining schedules, they are excluded from financial evaluations and the official Mineral Reserve Statement.

15.4. Mineral Reserve Estimate

Techno-economic project evaluation to a level of Prefeasibility study level has been completed for the Lofdal Project, following the 2019 CIM "Best Practice Guidelines for Estimation of Mineral Resources and Mineral Reserves." This assessment focused only on the Measured and Indicated sections of the Mineral Resources, resulting in a positive outcome that supports the estimation of Mineral Reserves as defined by the 2014 CIM "Definition Standards." Mineral Reserves are a subset of Mineral Resources and have been adjusted accordingly. Measured Mineral Resources were converted to Proven Mineral Reserves, while Indicated Mineral Resources became Probable Mineral Reserves. No Inferred Resources are included in the Mineral Reserves.

The Lofdal Rare Earths Project consists of two separate mineral deposits.

Area 4: (A4) Currently identified as the main mineralised area
Area 2B: (A2B) The secondary mineral deposit located approximately 2.5km WNW of A4

Table 63 Lofdal Mineral Reserves as of 01 December 2025

Reserve Category	Mineral Deposit	Tonnes	Rare Earths Grade			Contained Rare Earths Metal
			LREO	HREO	TREO	LREO	HREO	TREO
		(Mt)	(%)	(%)	(%)	(t)	(t)	(t)
Proven	Area 2B	-	-	-	-	-	-	-
	Area 4	6,19	0,068	0,144	0,211	4 194,0	8 893,2	13 087,1
Total Proven		6,19	0,068	0,144	0,211	4 194,0	8 893,2	13 087,1
Probable	Area 2B	1,90	0,075	0,094	0,169	1 430,3	1 792,8	3 223,1
	Area 4	23,91	0,076	0,091	0,167	18 269,3	21 761,6	40 030,7
Total Probable		25,81	0,076	0,091	0,168	19 699,7	23 554,4	43 253,8
Total Reserves		32,01	0,075	0,101	0,176	23 893,7	32 447,5	56 340,9

Notes on the Mineral Reserve:

Mineral reserves are reported at a cut-off grade of 0.100% TREO, based on a basket rare earths oxide price of USD 86.84/kg.
Reserves are based on open-pit mine designs with an average strip ratio of 6.8:1.
Metallurgical recovery at the hydrometallurgical plant is assumed at 62.2% for HREO and 52.55% for LREO.
Mass recovery for the >10mm primary crushed Low-Grade is assumed at 80%
Parameters at XRT Sorter is assumed at mass-pull of 25% and metal recovery of 65%
The reserve estimate was prepared by a Qualified Person (QP) in accordance with NI 43-101 and CIM Definition Standards (2014).
Mineral reserves are inclusive of mineral resources.

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16. Mining Methods

16.1. Overview

The proposed mining method is conventional open pit mining. Mineralised rock and waste would be drilled, blasted, loaded by hydraulic shovels and hydraulic excavators into off-highway dump trucks, and hauled to the processing plant.

The basis for the pit design work was the mineral resource block model that was developed by MSA as part of a NI 43--101-compliant mineral resource estimate (refer to Section 14).

There are two Lofdal deposits currently under consideration. These are to be mined as open pits, with the normal sequence of drilling, blasting and hauling. Due to the nature of the deposit, the resultant pits are relatively narrow along strike and deep. Currently no backfilling is contemplated.

The proposed mining method is the development of a slot-ramp along strike. This will enable selective waste mining on both sides of mineralised zones.

The target total ROM feed for processing is 3 010 000 tonnes/annum. On an annual basis, 1,91 Mtpa Low-Grade ore is fed to the XRT ore sorter after primary crushing and screening. 1,10 Mtpa High-Grade ore is fed to the crushing and milling circuit and then to the flotation and hydrometallurgical plant. The upgraded ore concentrate from the XRT sorter amounts to 0,38 Mtpa and joins the High-Grade ore stream before secondary crushing and milling.

Mineralised material with TREO >=0.05% and <0.10% is sent to a Marginal stockpile so it may be possible to process at a later stage if economics allow.

The aggregate Life of Mine for the two pits is estimated to be approximately 13 years, inclusive of pre-stripping activities. All cost and revenue modelling figures have been converted to US dollars.

16.2. Geotechnical Evaluation

NCMI commissioned SRK Consulting (South Africa) (Pty) Ltd (SRK) to conduct a geotechnical Feasibility Study (FS) for Lofdal. The study provides a targeted assessment to guide pit slope design for mining operations in A2B and A. Building on the 2023 PFS, this FS addresses data gaps—especially in footwall slopes—and integrates newly collected 2025 field data.

The 2025 programme included analysis of historical drilling, PFS data, and new FS data. Seven inclined drillholes (Figure 128) were logged and sampled, focusing on critical footwall slopes. Additional trench mapping, relogging, and lab tests contributed to an improved lithostructural and rock mass model, supporting refined slope stability analyses.

Rock mass at Lofdal is generally strong (Geological Strength Index 41-60), but local zones of weakness exist, particularly in albitite and alteration areas. Structural factors like foliation and faults significantly impact geotechnical conditions, raising risks of wedge and planar slope failures, especially in footwalls. While most domains meet stability criteria, mineralised or albitite zones present challenges. Crucially, recent footwall drillhole data was unusable for kinematic updates.

Groundwater monitoring after March 2025 rainfall showed substantial recharge, though the system remains poorly understood. Effective depressurisation and stormwater control are necessary for long-term stability, and simplified phreatic surfaces were used due to lack of pit-specific data.

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Figure 128 Historical 2023 and 2025 drillhole locations (SRK)

The study recommends further structural mapping, oriented core logging, and acoustic surveys to refine models, as well as ongoing groundwater and dewatering investigations. Slope design should be updated as new data arises, potentially allowing steeper slopes. Operational measures like controlled blasting, continuous monitoring, and stability radar use in Area 4 are also advised.

The 2025 FS advances understanding of Lofdal's geotechnical conditions and supports improved slope designs and project economics. Continued model refinement, especially regarding footwall slopes and groundwater, will be important for reliable mine development and project success.

The geotechnical design sectors (Figure 129) defined by SRK with the recommended slope angles and pit-wall configurations. The data was imported into MinePlan to inform the pit optimisation process and pit designs.

Figure 129 Geotechnical Design Sectors

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The slope angles and pit-wall configuration are summarised in Table 64 and Table 65. The recommended slope angles are exclusive of in pit roads/ramps. The inclusion of roads/ramps into the pits flattens the slope angles.

Table 64 A2B Domain Slope Design Recommendations

Slope Dip Range (°)	Domain	Batter Angle (°)	Barm Width (m)	Bench Height (m)	Stack Height (m)	Slope Height (m)	Geotechnical Berm Width (m)	Stack Angle (°)	Inter Ramp Angle (°)	Overall Slope Angle (°)
298 to 085	Area 2b: Domain A	80	4.3	10	50	127	15	63	59	54
085 to 240	Area 2b: Domain B	60	3.4	10	50	127	15	50	48	43
240 to 298	Area 2b: Domain C	65	2.9	10	50	127	15	55	53	47

Table 65 A4 Domain Slope Design Recommendations

Slope Dip Range (°)	Domain	Batter Angle (°)	Berm Width (m)	Bench Height (m)	Stack Height (m)	Slope Height (m)	Geotechnical Berm Width (m)	Stack Angle (°)	Inter Ramp Angle (°)	Overall Slope Angle (°)
080 to 180	Area 4: Domain A1	60	7.7	10	50	140	15	40	37	36
000 to 080	Area 4: Domain A2	75	3.8	10	50	150	15	60	57	52
154 to 180	Area 4: Domain B1	60	3.2	10	50	300	15	50	48	42
000 to 045	Area 4: Domain B2	90	5.7	10	50	270	15	63	57	50
154 to 210	Area 4: Domain C1	60	5.4	10	50	310	15	45	42	37
290 to 012	Area 4: Domain C2	90	5.5	10	50	330	15	63	57	50

The recommended slope angles were adjusted for the inclusion of the ramp widths of each sector. The adjusted slope angles were then used in the pit optimisations and finally in the pit designs. The adjusted slope angles and pit wall configurations are listed in Table 66 and Table 67.

Table 66 A2B Domain Slope Design inc. Ramps

MinePlan Slope Sector	Slope Sector/Domain	Bench Height	Batter Angle	Min. Berm Width	Max. Inter-stack Angle	Approx Crest Elev	Approx Toe Elev	Overall Slope Height	Number Stacks	Inter-stack Berm Widths	# Ramps	Ramp Width	Total Ramp Width	Calculated O’All Slope
(#)	(name)	(m)	(°)	(m)	(°)	(m)	(m)	(m)	(#)	(m)	#	m	m	(°)
1	Domain A	10,0	85°	4,30	63	950	820	130	2,0	30,0	0	20	0	53,2°
2	Domain B	10,0	60°	3,40	47	950	850	100	1,0	15,0	2	20	40	33,9°
3	Domain C	10,0	65°	2,90	53	950	850	100	1,0	15,0	0	20	0	47,8°

Table 67 A4 Domain Slope Design inc. Ramps

MinePlan Slope Sector	Slope Sector/Domain	Bench Height	Batter Angle	Min. Berm Width	Max. Inter-stack Angle	Approx Crest Elev	Approx Toe Elev	Overall Slope Height	Number Stacks	Inter-stack Berm Widths	# Ramps	Ramp Width	Total Ramp Width	Calculated O’All Slope
(#)	(name)	(m)	(°)	(m)	(°)	(m)	(m)	(m)	(#)	(m)	#	m	m	(°)
1	Domain A1	10,0	60°	7,70	37	980	840	140,0	2,00	30,0	1	15	15	30,0°
2	Domain A2	10,0	75°	3,80	57	980	830	150,0	2,00	30,0	0	15	0	49,7°
3	Domain B1	10,0	60°	3,20	48	980	680	300,0	5,00	75,0	2	20	40	38,0°
4	Domain B2	10,0	85°	5,70	57	980	710	270,0	5,00	75,0	3	20	60	40,8°
5	Domain C1	10,0	60°	5,40	42	980	670	310,0	6,00	90,0	2	20	40	33,1°
6	Domain C2	10,0	85°	5,50	57	980	650	330,0	6,00	90,0	3	20	60	42,5°

16.3. Hydrogeological Evaluation

SLR Environmental Consulting (Namibia) (Pty) Ltd (SLR) conducts quarterly groundwater monitoring at the Project site. The April 2025 report notes generally stable groundwater levels over time, with a slight decrease except for notable recharge from March 2025 rainfall (Figure 131). Borehole WW43415, nearest to Area 4, showed a groundwater level rise from

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7,13 m to 14,60 m. Due to limited data, a basic hard rock drawdown model was applied using pit depth and measurements from WW43415.

Figure 130 Groundwater monitoring boreholes at the project area (SLR)

Figure 131 GWL of quarterly monitored boreholes at Lofdal (SLR)

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16.4. Open Pit Optimisation

The pit optimisation for the two pits was accomplished using Hexagon's MinePlan Project Evaluator (MPPE) task suite. The input data and output results for MPPE is discussed below.

16.4.1. Summary Optimisation Parameters

The input optimisation parameters are summarised in

Table 68. This information has been collated the results from the PEA, additional resource drilling and updated resource, earlier cost estimation work done for the PFS, metallurgical test work, infrastructure plans, improved environmental- and social impact assessments, etc.

Table 68 Pit Optimisation Parameters Summary

Parameter	Units	a2b	a4
Macroeconomic Inputs
Exchange Rate	NAD:USD	18,5000	18,5000
Annual Discount Rate	%	8,00%	8,00%
Commodity Price
Base TREO Basket Price	USD/kg	86,8352	86,8352
Base HREO Basket Price	USD/kg	131,8303	131,8303
Base LREO Basket Price	USD/kg	45,2896	45,2896
Government Royalty	%	3,00%
Private Royalty	%	2,00%
Government Export Levy	%	0,00%
Selling Cost (Transport)	USD/t product	36,3100
Overall Slope Angle
All Sectors	deg	43,0° - 54,0°	36,0° - 52,0°
Mining Block Model Modifying Factors (0,05% TREO)
TREO Tonnage Factor	factor	0.928	0.970
TREO Grade Factor	factor	0.991	0.979
TREO Metal Factor	factor	0.920	0.949
Mining Operating Costs
Steady-state Ave Mining Cost	USD/t	2,9489	2,8499
Ref Ore Mining Cost	USD/t	2,8000	2,8000
Ref Waste Mining Cost	USD/t	2,8000	2,8000
Ore Depth Factor L&H	USD/t/5m-bench	0,0100	0,0100
Waste Depth Factor L&H	USD/t/5m-bench	0,0100	0,0100
Metallurgical Parameters
Process Recovery (TREO)	%	56,65%	56,38%
Milling Limit
High-Grade + Low-Grade	t/annum	3 010 000
Process Cost
Average Process Cost	USD/t ROM ore	28,8200	28,8200
High-Grade Process Cost	USD/t ROM ore	28,8200	28,8200
Low-Grade Process Cost	USD/t ROM ore	28,8200	28,8200
Incremental Ore Cost	USD/t ROM ore	0,2000	0,0000
Fixed Cost
Management - G&A	USD/annum	3 045 600
	USD/t ROM ore	1,0118

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16.4.2. Geological Block Models Input to Pit Optimisation

Variably sub-blocked Datamine geological block models were provided by the client as inputs. The dimension details of these as provided in Table 69 and Table 70.

Table 69 Area 2B a2bmod_20240215 Block Model

	min	max	size	extent	block count	sub-block size	sub-block count
x	466 900 0	467 850 0	10 00	950 0	95 0	2 00	5 00
y	7 754 300 0	7 755 250 0	10 00	950 0	95 0	2 00	5 00
z	600 0	1 000 0	5 00	400 0	80 0	1 00	5 00
					722 000,0

Table 70 Area 4 a2bmod_20240215 Block Model

	min	max	size	extent	block count	sub-block size	sub-block count
x	469 500,0	471 400,0	10,00	1 900,0	190,0	2,00	5,00
y	7 752 800,0	7 754 300,0	10,00	1 500,0	150,0	2,00	5,00
z	500,0	1 200,0	5,00	700,0	140,0	1,00	5,00
					3 990 000,0

The Datamine block models were imported in MinePlan and verified against unconstrained statistics reported by MSA. The MinePlan sub-blocked models are named "a2sbm15.dat" and "a4sbm15.dat" for A2B and A4 respectively.

Two selective mining unit (SMU) block sizes were evaluated to regularise the sub-blocked models. This process was undertaken to simulate mining loss and dilution for the purposes of mine planning. These SMUs are $10,0\mathrm{m} \times 10,0\mathrm{m} \times 5,0\mathrm{m}$ ( $x, y, z$ ) for $500,0\mathrm{m}^3$ ( $\approx 1400$ tonnes) and $5,0\mathrm{m} \times 5,0\mathrm{m} \times 2,5\mathrm{m}$ for $62,5\mathrm{m}^3$ ( $\approx 175$ tonnes). Since only the M+I resources are considered for PFS level accuracy, the decision on the SMU factors was based on M+I resources. The results of the two regularised models as shown in Table 69 and Table 70. The regularised models are named "a2breg.dat" and "a2breg2.dat" for A2B and "a4reg.dat" and "a4reg2.dat" for A4.

The grade-tonnage curves for M+I resources of the two pairs of the regularised models with the original sub-blocked models are shown in Figure 132. The grade-tonnage curves show the dilution- and loss impact of regularisation on tonnage and grade. The modifying factors for a cut-off of $0,05\%$ TREO is illustrated in Table 71. For mine planning work, it was decided to proceed with the $5,0\mathrm{m} \times 5,0\mathrm{m} \times 2,5\mathrm{m}$ SMU size with the following factors:

A2B: Net tonnage loss of $7.2\%$ , grade reduction of $0.9\%$ for a TREO metal loss of $8.0\%$ .
A4: Net tonnage loss of $3.0\%$ , grade reduction of $2.1\%$ for a TREO metal loss of $5.1\%$ .

Table 71 M+I Modifying Factors for Regularised Models

Resource Model	Regularised Model	Modifying Factors ≥ 0,05% TREO
		Tonnes	TREO %	TREO Metal
A2B	a2breg.dat	0,869	0,980	0,851
	a2breg2.dat	0,928	0,991	0,920
A4	a4reg.dat	0,940	0,963	0,905
	a4reg2.dat	0,970	0,979	0,949

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Table 72 Area 2B Sub-blocked Model vs. Regularised Models

Sub-blocked Model
MZone	Volume	Tonnes	DENS	TREO %	HREO %	LREO %	La2O3	Ce2O3	Pr2O3	Nd2O3	Sm2O3	Eu2O3	Gd2O3	Tb2O3	Dy2O3	Ho2O3	Er2O3	Tm2O3	Yb2O3	Lu2O3	Y2O3
MZONE_00	305 164 452	863 615 399	2,83	0,001	0,000	0,001	1,207	2,497	0,558	2,837	0,461	0,167	0,588	0,079	0,452	0,090	0,268	0,041	0,261	0,042	2,757
MZONE_01	5 508 344	15 286 753	2,78	0,099	0,063	0,035	78,449	138,233	16,419	78,035	41,969	15,455	53,070	9,898	62,596	12,557	36,589	5,397	34,385	5,017	399,791
MZONE_02	2 337 820	6 498 500	2,78	0,100	0,044	0,056	141,223	239,958	26,667	110,461	42,977	13,586	47,705	8,365	47,715	8,673	23,308	3,225	20,021	2,920	259,619
MZONE_03	2 248 736	6 322 600	2,81	0,064	0,036	0,028	52,999	107,041	13,939	73,075	34,825	10,783	37,904	6,053	37,077	7,003	19,869	2,762	17,337	2,515	218,812
MZONE_04	25 324	71 970	2,84	0,056	0,031	0,025	59,585	109,539	12,019	48,910	18,570	7,089	24,353	4,867	32,015	6,311	17,640	2,513	16,073	2,363	201,085
MZONE_05	42 088	118 133	2,81	0,051	0,025	0,026	65,774	117,819	12,611	46,983	15,457	5,962	22,327	4,366	26,796	5,070	13,632	1,879	7,189	1,918	157,970
MZONE_06	199 652	557 572	2,79	0,116	0,036	0,080	222,837	353,798	36,790	139,082	42,671	13,532	43,017	6,623	37,822	6,864	19,768	2,527	15,989	2,378	211,549
MZONE_07	171 244	474 006	2,77	0,075	0,029	0,047	102,067	194,327	23,431	107,098	41,249	11,637	37,275	5,569	29,515	5,240	14,198	1,938	12,153	1,777	167,489
MZONE_08	1 049 588	2 916 311	2,78	0,116	0,031	0,085	232,625	377,968	40,821	156,491	40,109	12,600	41,379	6,185	33,544	5,834	15,148	2,032	12,484	1,826	178,712
MZONE_09	61 168	169 424	2,77	0,080	0,031	0,049	144,362	228,397	23,263	80,681	16,307	7,083	25,654	5,460	34,371	6,608	18,319	2,492	14,953	2,294	189,666
Model Total	316 808 416	896 030 670	2,83	0,005	0,002	0,002	4,890	8,881	1,285	6,061	1,903	0,657	2,276	0,378	2,272	0,442	1,268	0,185	1,168	0,174	13,779
Model Waste	305 164 452	863 615 399	2,83	0,001	0,000	0,001	1,207	2,497	0,558	2,837	0,461	0,167	0,588	0,079	0,452	0,090	0,268	0,041	0,261	0,042	2,757
Model Ore	11 643 964	32 415 271	2,78	0,093	0,050	0,043	103,026	178,972	20,650	91,938	40,329	13,727	47,262	8,332	50,766	9,813	27,898	4,010	25,327	3,697	307,444
Regularized 10,0 x 10,0m x 5,0m Model
	Volume	Tonnes	DENS	TREO %	HREO %	LREO %	La2O3	Ce2O3	Pr2O3	Nd2O3	Sm2O3	Eu2O3	Gd2O3	Tb2O3	Dy2O3	Ho2O3	Er2O3	Tm2O3	Yb2O3	Lu2O3	Y2O3
Model Total	316,213,113	894,349,496	2.83	0.005	0.002	0.002	4,884	8,871	1,284	6,056	1,901	0,657	2,274	0,378	2,270	0,442	1,267	0,185	1,167	0,174	13,769
Model Waste	304,801,472	862,525,840	2.83	0.002	0.001	0.001	1,703	3,351	0,652	3,236	0,625	0,223	0,783	0,115	0,669	0,132	0,388	0,059	0,368	0,057	4,082
Model Ore	11,411,640	31,823,656	2.79	0.084	0.045	0.039	91,105	158,500	18,411	82,496	36,473	12,416	42,701	7,510	45,676	8,824	25,089	3,610	22,831	3,331	276,328
Regularized 5,0 x 5,0m x 2,5m Model
	Volume	Tonnes	DENS	TREO %	HREO %	LREO %	La2O3	Ce2O3	Pr2O3	Nd2O3	Sm2O3	Eu2O3	Gd2O3	Tb2O3	Dy2O3	Ho2O3	Er2O3	Tm2O3	Yb2O3	Lu2O3	Y2O3
Model Total	316 811 750	896 040 600	2,83	0,005	0,002	0,002	4,891	8,881	1,285	6,061	1,903	0,657	2,276	0,378	2,272	0,442	1,268	0,185	1,168	0,174	13,779
Model Waste	305 221 375	863 742 781	2,83	0,001	0,001	0,001	1,471	2,951	0,608	3,049	0,548	0,197	0,692	0,098	0,568	0,113	0,332	0,051	0,318	0,050	3,462
Model Ore	11 590 375	32 297 819	2,79	0,088	0,047	0,041	96,342	167,476	19,388	86,590	38,125	12,976	44,649	7,862	47,859	9,249	26,296	3,782	23,902	3,488	289,696

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Table 73 Area 4 Sub-blocked Model vs. Regularised Models

Sub-blocked Model

MZone	Volume	Tonnes	DENS	TREO %	HREO %	LREO %	La2O3	Ce2O3	Pr2O3	Nd2O3	Sm2O3	Eu2O3	Gd2O3	Tb2O3	Dy2O3	Ho2O3	Er2O3	Tm2O3	Yb2O3	Lu2O3	Y2O3
MZONE_00	1 276 239 228	3 611 757 015	2,83	0,001	0,000	0,001	0,970	1,935	0,765	3,619	0,337	0,087	0,311	0,054	0,299	0,060	0,193	0,028	0,190	0,028	1,903
MZONE_01	32 019 352	87 659 330	2,74	0,122	0,062	0,060	151,165	275,792	29,838	110,121	33,874	11,947	47,118	9,271	59,740	12,319	35,820	5,250	32,348	4,709	401,831
MZONE_02	4 701 472	12 937 637	2,75	0,098	0,036	0,062	157,955	288,398	31,236	113,916	27,846	8,981	32,363	5,717	35,085	7,053	20,258	2,841	17,731	2,559	228,304
MZONE_03	2 719 984	7 466 426	2,75	0,116	0,041	0,075	207,514	353,877	36,281	125,998	26,646	8,617	32,216	6,181	39,371	8,085	23,832	3,523	22,051	3,202	259,266
MZONE_04	1 408 016	3 959 496	2,81	0,078	0,029	0,049	128,159	228,240	24,005	88,811	21,169	7,004	25,433	4,640	29,227	5,844	16,725	2,337	14,420	2,050	186,179
MZONE_05	667 624	1 825 303	2,73	0,142	0,091	0,051	126,303	233,272	25,669	95,886	25,948	10,186	50,294	11,999	88,093	19,231	56,595	7,871	49,610	7,317	608,565
MZONE_06	397 944	1 134 251	2,85	0,097	0,026	0,071	192,651	336,925	35,183	123,461	24,329	6,936	23,406	4,001	25,583	5,047	14,291	1,984	12,616	1,854	163,256
MZONE_07	1 817 648	5 066 754	2,79	0,145	0,037	0,108	313,840	511,898	50,310	171,873	32,891	10,007	34,382	5,895	36,568	7,201	20,594	2,888	18,015	2,570	230,471
MZONE_08	1 256 992	3 481 028	2,77	0,088	0,043	0,044	112,255	204,158	21,914	81,816	21,744	7,456	31,463	6,185	41,683	8,693	25,902	3,804	23,970	3,513	280,986
MZONE_09	1 769 380	4 990 649	2,82	0,093	0,024	0,069	180,953	321,323	34,434	127,994	27,568	7,990	25,570	4,075	24,077	4,579	12,726	1,783	11,300	1,670	147,870
MZONE_10	210 964	590 765	2,80	0,129	0,042	0,087	225,589	410,861	43,547	155,242	34,808	12,233	40,965	6,881	41,560	8,175	22,790	3,180	19,950	2,923	260,437
MZONE_11	726 928	1 961 419	2,70	0,100	0,040	0,059	144,627	272,005	29,879	112,161	32,814	10,312	38,560	6,987	41,239	8,084	21,978	3,152	19,144	2,795	251,437
MZONE_12	2 354 616	6 378 484	2,71	0,115	0,051	0,063	152,614	287,827	32,055	120,664	39,380	13,003	50,768	9,199	54,584	10,660	28,418	4,002	23,800	3,449	315,843
MZONE_13	693 752	1 848 596	2,66	0,067	0,026	0,041	101,866	189,265	20,736	76,070	20,327	7,321	25,573	4,374	26,416	5,146	14,522	2,060	13,454	2,044	157,248
MZONE_14	753 104	2 037 317	2,71	0,054	0,022	0,032	74,571	147,763	16,816	64,014	17,704	6,200	21,075	3,659	21,761	4,290	12,364	1,855	12,831	2,100	133,956
MZONE_15	353 728	976 289	2,76	0,058	0,022	0,037	91,798	162,687	18,005	69,191	23,993	6,954	24,264	3,552	21,582	4,156	11,723	1,673	10,774	1,601	129,040
Model Total	1 328 090 732	3 754 070 760	2,83	0,005	0,002	0,003	6,928	12,674	1,896	7,730	1,523	0,494	1,887	0,357	2,235	0,456	1,338	0,194	1,218	0,178	14,719
Model Waste	1 276 239 228	3 611 757 015	2,83	0,001	0,000	0,001	0,970	1,935	0,765	3,619	0,337	0,087	0,311	0,054	0,299	0,060	0,193	0,028	0,190	0,028	1,903
Model Ore	51 851 504	142 313 745	2,74	0,115	0,053	0,062	158,135	285,219	30,597	112,068	31,623	10,834	41,862	8,044	51,355	10,512	30,391	4,418	27,302	3,975	339,967

Regularized 10,0 x 10,0m x 5,0m Model

	Volume	Tonnes	DENS	TREO %	HREO %	LREO %	La2O3	Ce2O3	Pr2O3	Nd2O3	Sm2O3	Eu2O3	Gd2O3	Tb2O3	Dy2O3	Ho2O3	Er2O3	Tm2O3	Yb2O3	Lu2O3	Y2O3
Model Total	1 328 126 000	3 754 174 107	2,83	0,005	0,002	0,003	6,928	12,671	1,896	7,729	1,522	0,494	1,886	0,357	2,234	0,456	1,338	0,194	1,217	0,178	14,717
Model Waste	1 276 848 500	3 613 113 089	2,83	0,001	0,000	0,001	1,596	3,041	0,878	4,026	0,438	0,119	0,429	0,076	0,432	0,087	0,269	0,039	0,257	0,038	2,745
Model Ore	51 277 500	141 061 018	2,75	0,106	0,050	0,056	143,465	259,322	27,960	102,590	29,306	10,108	39,205	7,568	48,395	9,922	28,710	4,179	25,812	3,758	321,363

Regularized 5,0 x 5,0m x 2,5m Model

	Volume	Tonnes	DENS	TREO %	HREO %	LREO %	La2O3	Ce2O3	Pr2O3	Nd2O3	Sm2O3	Eu2O3	Gd2O3	Tb2O3	Dy2O3	Ho2O3	Er2O3	Tm2O3	Yb2O3	Lu2O3	Y2O3
Model Total	1 328 040 813	3 753 930 753	2,83	0,005	0,002	0,003	6,928	12,673	1,896	7,730	1,523	0,494	1,887	0,357	2,235	0,456	1,338	0,194	1,218	0,178	14,719
Model Waste	1 276 359 750	3 611 900 455	2,83	0,001	0,000	0,001	1,303	2,524	0,826	3,836	0,391	0,104	0,375	0,066	0,371	0,074	0,234	0,034	0,226	0,033	2,359
Model Ore	51 681 063	142 030 298	2,75	0,110	0,051	0,059	149,972	270,786	29,124	106,765	30,307	10,418	40,329	7,767	49,627	10,167	29,405	4,277	26,427	3,848	329,043

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Figure 132 Grade-Tonnage Curves for Regularised Models vs. Sub-blocked Models

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16.4.3. Processing Plant Capacity

The processing plant capacity is based on a flowsheet that separates the ROM feed into High-Grade (HG) and Low-Grade (LG) streams as illustrated in Figure 133. Total ROM feed is set at 3 010 000 tonnes/annum. This is made up of 1 100 000 tonnes of High-Grade and 1 910 000 tonnes of Low-Grade. Refer to Section 17, Recovery Methods for more detail.

Figure 133 Process Plant Flowsheet

16.4.4. General & Administration (Fixed) Costs

The G&A overheads costs are estimated at USD 3,045M per annum at full production. This is factored to USD 1,012/tonne ROM feed and was applied to the MPPE Economics Task in the pit optimisation model. (Table 74)

Table 74 General and Administration (Fixed) Costs

Item	Unit	Value
G&A and Fixed Costs	USD/tonne processed	1,012
Total Site G&A	USD/annum	3 045 600
Total Process Feed (HG + LG)	tonnes processed/annum	3 010 000

16.4.5. Processing Recovery

The process recoveries for the 15 different REO are listed in Table 75 and is quoted as net-of-ROM REO recoveries. This important for the LG feed that is going through a screening process to remove the fines (<10mm) from the ROM feed before sending the ≥10mm fraction

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to the sorter feed. The XRT mass pull and REO-recovery is based on test work set at 25% and 65% TREO respectively.

Table 75 Net-of-ROM Process Recovery
| REO Item | Unit | HG Recovery | LG Recovery |
| --- | --- | --- | --- |
| Average TREO Net-of-ROM Process Recovery | % | 59,38% | 30,88% |
| HREO Net-of-ROM Process Recovery | % | 62,24% | 32,37% |
| LREO Net-of-ROM Process Recovery | % | 52,55% | 27,32% |
| Contributing RE Oxides OM Process Recovery | | | |
| La2O3 Lanthanum(III) oxide (Lights) | % | 52,5% | 27,3% |
| Ce2O3 Cerium(III) oxide (Lights) | % | 52,5% | 27,3% |
| Pr2O3 Praseodymium(III) oxide (Lights) | % | 52,5% | 27,3% |
| Nd2O3 Neodymium(III) oxide (Lights) | % | 52,5% | 27,3% |
| Sm2O3 Samarium(III) oxide (Lights) | % | 52,5% | 27,3% |
| Eu2O3 Europium(III) oxide (Heavies) | % | 62,2% | 32,4% |
| Gd2O3 Gadolinium(III) oxide (Heavies) | % | 62,2% | 32,4% |
| Tb2O3 Terbium(III) oxide (Heavies) | % | 62,2% | 32,4% |
| Dy2O3 Dysprosium(III) oxide (Heavies) | % | 62,2% | 32,4% |
| Ho2O3 Holmium (III) oxide (Heavies) | % | 62,2% | 32,4% |
| Er2O3 Erbium(III) oxide (Heavies) | % | 62,2% | 32,4% |
| Tm2O3 Thulium(III) oxide (Heavies) | % | 62,2% | 32,4% |
| Yb2O3 Ytterbium(III) oxide (Heavies) | % | 62,2% | 32,4% |
| Lu2O3 Lutetium(III) oxide (Heavies) | % | 62,2% | 32,4% |
| Y2O3 Yttrium oxide (Heavies) | % | 62,2% | 32,4% |

16.4.6. Mining Costs

The mining methods proposed for this project are founded on established conventional open pit mining techniques. The planned operations will comprise drilling and blasting activities to fragment the rock, after which both mineralised material and waste will be systematically loaded and hauled using standard mining equipment. This methodology is intended to facilitate the efficient extraction of resources while upholding recognized industry standards for safety and productivity.

The mining costs applied in this analysis were derived from a combination of data sources, including results from previous Preliminary Feasibility Studies (PFS) and proposals received from mining contractors engaged in projects of a similar nature. By referencing established cost parameters and contractor insights from comparable mining operations, the estimates provide a robust basis for evaluating the financial requirements of the current project and ensure that the projected costs are grounded in industry practice and recent market conditions.

The summary mining costs applied to the pit optimisation listed in Table 76. Overall mining cost is estimated at ≈USD 2,85/tonne mined for A4 and ≈USD 2,95/tonne mined for A2B due to the longer haul to the plant and narrow, confined pit benches.

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Table 76 Mining Costs
| Cost Item | Unit | Area 2B | Area 4 |
| --- | --- | --- | --- |
| Estimated Total Mining Cost | USD/tonne mined | 2,9489 | 2,8499 |
| Ore Mining Cost @ pit exit | USD/tonne ore mined | 2,5841 | 2,4361 |
| Waste Mining Cost @ pit exit | USD/tonne waste mined | 2,5841 | 2,4361 |
| Loading Cost | USD/tonne mined | 0,4125 | 0,4125 |
| Hauling Cost @ Pit Exit | USD/tonne mined | 0,9628 | 0,8148 |
| Drilling & Blasting | USD/tonne mined | 0,5936 | 0,5936 |
| Support, Rehandle & Other Mining Cost | USD/tonne mined | 0,2210 | 0,2210 |
| Mining Contractor Fixed Cost | USD/tonne mined | 0,2702 | 0,2702 |
| Owner's Mining Team Cost | USD/tonne mined | 0,1240 | 0,1240 |
| Incremental Ore Hauling Cost/5m | USD/tonne mined/5m | 0,0224 | 0,0187 |
| Incremental Waste Hauling Cost/5m | USD/tonne mined/5m | 0,0176 | 0,0152 |

16.4.7. Taxes, Levies and Selling Costs

16.4.7.1. Selling Costs

The cost of transporting the concentrate to Walvis Bay is estimated at a cost of USD 36,31/tonne concentrate, as per the PEA. The metal separation cost is excluded from the pit optimisation.

Table 77 Selling Cost
| Selling Cost Item | Units | Value | Basis / Comments |
| --- | --- | --- | --- |
| Selling Cost (Logistics : Insurance) | USD/tonne concentrate | 36,3100 | Confirmed from the PEA |
| Sales Agent Commission | % of Revenue | 0,00% | Not applicable |

16.4.7.2. Royalties

The applicable royalties are a Namibian government royalty of 3% on revenue and 2% Landowner royalty payable on revenue. (Table 78)

Table 78 Royalties, Taxes and Levies
| Royalties, Taxes, Levies | Unit | Value | Basis / Comments |
| --- | --- | --- | --- |
| Government of Namibia Royalty | %¹ | 3,00% | GRN Notice No. 45, 18 March 2009 |
| Private Royalty | %¹ | 2,00% | NCMI - 07-May-2024 |
| Export Levy | %¹ | 0,00% | 0% GRN Export Levy on RE products |
| Namibia Mining Company Tax Rate | % | 37,50% | Income Tax Act, 1981 (Act 24 of 1981); Income Tax Amendment Act, 2011 (Act No. 3 of 2011) |
| ¹ – Percentage of Revenue | | | |

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16.4.8. Processing Costs

The applied processing cost of USD 30.55/tonne MILLED is modelled from the updated flowsheet, including the LG processed via the XRT sorter. The processing cost for the pit optimisation is expressed as USD on a per tonne ROM basis. (Table 79)

Table 79 Processing Costs

Item	Unit	High Grade	Low Grade	Total Feed
Total ROM Feed	tonnes/annum	1 100 000	1 910 000	3 010 000
Total MILL Feed	tonnes/annum	1 100 000	382 000	1 482 000
Total Processing Cost	USD/tonne ROM processed	28,8200	6,0790	14,3897
Total Processing Cost	USD/tonne MILL Feed	28,8200	30,3950	30,5550

16.4.9. Open Pit Constraints and Mining Limits

There are no pit constraints applicable for either of the two pit areas in the pit shell generation. During the Best/Worst Case Analysis, the ROM mining rates for A2B was set at 0,5Mtpa and 3,01 Mtpa for A4.

16.4.10. Mining Recovery and Dilution

Mining recovery and dilution is discussed in Section 16.4.2 Geological Block Models Input to Pit Optimisation as follows:

A2B: Net tonnage loss of 7,2%, grade reduction of 0,9% for a TREO metal loss of 8,0%.
A4: Net tonnage loss of 3,0%, grade reduction of 2,1% for a TREO metal loss of 5,1%.

16.4.11. Product Prices

The product prices for individual rare earth oxides were applied rather than a basket price. The prices Table 80 are provided by NCMI for the pit optimisation. All the rare earth oxide grade values are carried through to the optimisation and directly applied to each grade value.

Table 80 Rare Earth Product Prices

Rare Earth Oxide	Class	Gross Price	GRN Royalty	Private Royalty	Transport Cost	Net Price
		USD/kg	USD/kg	USD/kg	USD/kg	USD/kg
TREO Total rare earth oxides	TREO	86.84	2.61	1.74	0.04	82.46
HREO Heavy rare earth oxides	HREO	131.83	3.95	2.64	0.04	125.20
LREO Light rare earth oxides	LREO	45.29	1.36	0.91	0.04	42.99

16.4.12. Net Revenue

The "Net Prices" in Table 80 were applied to the pit optimisation as the royalties are "costs" based on revenue that each rare earth oxide attract.

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16.5. Optimisation Results

The pit optimisations for A2B and A4 were executed independently. A sequence of 31 nested pits was developed using revenue factors (RF) ranging from RF=0,60 to RF=1,20 in increments of 0,20. Employing the regularised block models detailed in Section 16.4.2, measured and indicated resources (M+I) were incorporated into the optimisation process, while inferred material was classified as waste. The materials were further categorised into five types (GTYPE) according to TREO% grade, as outlined below:

Table 81 Pit Optimisation Material Types

	GTYPE CODE	Grade Type	TREO% Grade Range	Pit Optimisation Destination
GTYPE	1	Waste	TREO% < 10,00%	Waste Dump
GTYPE	2	Mineralised Waste	0,06% ≤ TREO% < 0,08%	Waste Dump
GTYPE	3	Marginal	0,08% ≤ TREO% < 0,10%	Marginal Stockpile
GTYPE	4	Low Grade	0,10% ≤ TREO% < 0,16%	Low-Grade ROM
GTYPE	5	High Grade	TREO% ≥ 0,16%	High-Grade ROM

The MPPE configuration for A4 is depicted in Figure 134. The Block Model Source task establishes connections to the resource model and identifies essential attributes such as materials, grade, cost, and slope for further analysis. These block model attributes are subsequently transferred to the Economics Model task. Here, each block is assigned a monetary value that reflects its mining and processing costs as well as its potential for value generation; in this instance, this process is performed across 31 economic models.

The economic models are then advanced to the Pit Shell Generation task, where they are processed using either the Pseudoflow or Lerchs-Grossmann algorithm. This step determines the optimal pit boundaries for various economic models while adhering to the required slope constraints.

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Figure 134 A4 Project Evaluator Task Setup

A Best-Worst Case analysis was performed for both pits, and the outcomes were thoroughly examined. The detailed results and conclusions are presented in the subsequent discussion.

16.5.1. Pit A2B

The Best-Worst case analysis for Pit A2B was performed using a ROM (HG + LG) feed rate of 500 ktpa, and the results follow the typical pattern seen in such studies. The summary of findings is presented in Table 4 and illustrated in Figure 135.

There are a two prominent of inflexion points along the value curves (Figure 135) where there are sudden increases in values – at pit shells RF=0.72 and RF=0.84. These are ignored in the selection of interim pushbacks as the optimum pit shell is quite small and cannot accommodate an interim pushback.

Table 82 A2B Best-Worst Case Analysis

	Tonnes in Pit Shell				Worst Case	Best Case	Average Case
Pit Shell Rev. Factor	High-Grade Ore	Low-Grade Ore	Total Mining	Net Value	Net Present Value	Net Present Value	Net Present Value	Strip Ratio
RF-0.60	535 255	488 079	4 703 767	46 870 380	43 038 965	43 038 965	43 038 965	10,99
RF-0.62	561 107	530 253	5 268 586	48 949 450	44 562 268	44 754 145	44 658 207	10,14
RF-0.64	590 736	581 430	6 085 368	50 986 330	46 019 653	46 434 520	46 227 087	9,05
RF-0.66	602 345	639 127	6 709 330	52 226 083	46 913 220	47 457 284	47 185 252	8,28
RF-0.68	602 858	642 731	6 743 118	52 295 770	46 965 873	47 514 774	47 240 324	8,24
RF-0.70	636 153	692 446	7 773 252	54 169 440	48 287 031	49 060 504	48 673 768	7,43
RF-0.72	695 072	762 867	9 449 043	56 765 773	50 148 476	51 202 413	50 675 445	6,51
RF-0.74	698 905	769 283	9 563 782	56 971 454	50 285 065	51 372 095	50 828 580	6,46
RF-0.76	699 080	770 154	9 574 877	56 983 698	50 292 839	51 382 196	50 837 518	6,45
RF-0.78	715 437	793 698	10 215 599	57 593 669	50 647 916	51 836 052	51 241 984	6,16

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	Tonnes in Pit Shell				Worst Case	Best Case	Average Case
Pit Shell Rev. Factor	High-Grade Ore	Low-Grade Ore	Total Mining	Net Value	Net Present Value	Net Present Value	Net Present Value	Strip Ratio
RF-0.80	716 641	795 066	10 276 061	57 639 020	50 657 722	51 870 694	51 264 208	6,13
RF-0.82	716 641	797 305	10 294 865	57 652 354	50 655 083	51 880 880	51 267 982	6,12
RF-0.84	910 847	956 528	16 108 385	62 054 817	52 185 999	55 243 769	53 714 884	4,66
RF-0.86	937 969	979 378	17 132 517	62 643 223	52 348 188	55 693 232	54 020 710	4,51
RF-0.88	937 969	987 678	17 215 298	62 666 423	52 347 285	55 710 954	54 029 120	4,48
RF-0.90	949 042	1 005 902	17 551 774	62 961 256	52 476 569	55 936 166	54 206 368	4,43
RF-0.92	955 217	1 015 496	17 837 218	63 031 163	52 446 832	55 989 566	54 218 199	4,39
RF-0.94	962 985	1 065 214	18 616 775	63 127 095	52 182 580	56 062 845	54 122 713	4,21
RF-0.96	965 031	1 071 560	18 731 526	63 139 819	52 119 036	56 059 847	54 089 442	4,19
RF-0.98	969 344	1 083 811	18 938 790	63 241 668	52 057 610	56 131 884	54 094 747	4,16
RF-1.00	969 344	1 084 500	18 944 656	63 240 458	52 052 401	56 131 028	54 091 715	4,16
RF-1.02	976 091	1 103 016	19 251 077	63 463 448	52 042 184	56 288 745	54 165 465	4,11
RF-1.04	977 481	1 107 675	19 341 621	63 443 583	51 982 434	56 274 695	54 128 565	4,09
RF-1.06	977 481	1 108 023	19 345 324	63 441 882	51 979 092	56 273 492	54 126 292	4,09
RF-1.08	990 408	1 128 639	20 037 467	63 267 434	51 492 717	56 150 108	53 821 413	3,98
RF-1.10	990 408	1 129 691	20 050 539	63 260 431	51 480 594	56 145 155	53 812 875	3,98
RF-1.12	990 579	1 131 566	20 078 429	63 249 971	51 457 145	56 137 757	53 797 451	3,97
RF-1.14	990 748	1 132 432	20 096 437	63 239 099	51 441 843	56 130 067	53 785 955	3,97
RF-1.16	990 921	1 132 783	20 107 016	63 231 573	51 432 003	56 124 744	53 778 374	3,97
RF-1.18	999 746	1 166 392	20 580 502	63 396 083	51 280 479	56 241 099	53 760 789	3,89
RF-1.20	999 919	1 169 521	20 612 185	63 388 227	51 263 974	56 235 543	53 749 759	3,89

Figure 135 A2B Best/Worst Case Analysis

Figure 136 presents the TREO% grades and strip ratios for the incremental pit shells. It is observed that the strip ratio decreases significantly as the pit shells expand up to RF=0.84 pit shell, at which point there is a notable increase in the TREO% grade. The A2B pit shell, being relatively small, would likely be mined in a single pushback. Accordingly, it is advisable to determine the optimal pit shell by considering the Worst-Case analysis.

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Figure 136 shows the TREO% grades and strip ratio for the incremental pit shells. It is notable that the strip ratio steeply decreases with increasing pit shells up to RF=0,84 pit shell. Also is also a notable sharp increase in TREO% grade at that point.

Figure 136 A2B Pit Shell Strip Ratio and Grade

16.5.2. Pit A4

The Best-Worst case analysis for Pit A4 was performed using a ROM (HG + LG) feed rate of 3,010 ktpa, and the results seem somewhat different from the typical pattern seen in such studies. The summary of findings is presented in Table 83 and illustrated in Figure 137.

The Best-Case value curve in Figure 137 is as one would expect; showing some mild inflexion points and then starts to plateau close to the RF-1,00 pit shell before decreasing slowly beyond the RF=1,00 pit shell. The last notable inflexion point on this value curve occurs at pit RF=0,92 pit shell. The Worst-Case value curve is somewhat "erratic" with several significant peaks and troughs with the highest peak realised at RF= 0,78 pit shell at a NPV value of USD 926,6M. The Average-Case value curve mirrors the Worst-Case pattern at a higher value.

Table 83 Pit A4 Best-Worst Case Analysis

	Tonnes in Pit Shell				Worst Case	Best Case	Average Case
Pit Shell Rev. Factor	High-Grade Ore	Low-Grade Ore	Total Mining	Net Value	Net Present Value	Net Present Value	Net Present Value	Strip Ratio
RF-0.60	6 722 991	6 976 533	69 001 644	987 918 190	846 300 156	832 383 051	839 341 604	4,04
RF-0.62	6 870 421	7 327 940	75 950 339	986 759 286	800 278 318	830 975 716	815 627 017	4,35
RF-0.64	6 914 413	7 411 783	76 587 999	989 043 085	800 803 885	832 591 009	816 697 447	4,35
RF-0.66	7 602 549	9 695 148	93 414 010	1 035 497 345	806 574 976	864 415 948	835 495 462	4,40
RF-0.68	7 750 783	10 126 672	96 975 534	1 044 812 006	849 418 801	870 516 041	859 967 421	4,42
RF-0.70	7 900 845	10 427 933	99 821 245	1 053 791 317	855 218 928	876 396 515	865 807 722	4,45

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	Tonnes in Pit Shell				Worst Case	Best Case	Average Case
Pit Shell Rev. Factor	High-Grade Ore	Low-Grade Ore	Total Mining	Net Value	Net Present Value	Net Present Value	Net Present Value	Strip Ratio
RF-0.72	8 007 494	10 750 716	102 666 238	1 059 826 545	810 453 901	880 348 935	845 401 418	4,47
RF-0.74	8 037 049	10 852 096	103 462 347	1 061 351 566	810 013 799	881 347 658	845 680 729	4,48
RF-0.76	8 121 517	11 051 444	105 358 399	1 065 401 572	881 301 398	883 999 972	882 650 685	4,50
RF-0.78	8 279 650	11 502 884	109 817 527	1 072 223 070	926 649 372	888 126 541	907 387 957	4,55
RF-0.80	8 588 674	12 469 206	122 118 717	1 085 514 617	851 887 875	896 215 435	874 051 655	4,80
RF-0.82	8 692 682	12 781 884	125 714 727	1 089 235 363	853 162 042	898 471 625	875 816 834	4,85
RF-0.84	8 896 667	13 283 345	132 243 666	1 096 678 825	856 076 712	902 385 165	879 230 939	4,96
RF-0.86	9 213 747	14 204 607	143 457 725	1 105 947 634	798 287 477	907 589 267	852 938 372	5,13
RF-0.88	9 334 807	14 673 652	148 069 610	1 109 570 552	811 661 445	909 623 404	860 642 425	5,17
RF-0.90	9 427 096	15 046 108	151 809 479	1 111 999 553	810 747 571	910 987 201	860 867 386	5,20
RF-0.92	9 845 804	16 401 234	166 902 428	1 126 705 316	780 206 920	918 336 354	849 271 637	5,36
RF-0.94	9 920 884	16 648 511	169 768 420	1 128 272 217	778 691 700	919 150 945	848 921 323	5,39
RF-0.96	10 120 788	17 233 958	177 618 305	1 130 368 358	774 067 959	920 240 674	847 154 317	5,49
RF-0.98	11 633 614	20 973 488	232 622 093	1 136 799 650	840 247 839	920 021 911	880 134 875	6,13
RF-1.00	11 792 598	21 688 353	239 849 064	1 137 774 423	834 863 806	920 456 375	877 660 091	6,16
RF-1.02	12 038 236	22 383 354	249 138 969	1 137 295 844	827 589 334	920 243 068	873 916 201	6,24
RF-1.04	12 562 630	23 543 423	269 046 631	1 134 585 178	813 585 492	919 034 903	866 310 198	6,45
RF-1.06	12 587 643	23 697 485	270 475 901	1 134 237 300	812 555 964	918 830 790	865 693 377	6,45
RF-1.08	12 663 761	23 959 327	274 211 241	1 133 473 965	825 742 427	918 515 767	872 129 097	6,49
RF-1.10	12 725 024	24 224 288	277 406 718	1 132 724 103	834 294 331	918 206 304	876 250 318	6,51
RF-1.12	12 849 367	24 578 027	282 645 939	1 131 332 855	832 155 201	917 632 146	874 893 674	6,55
RF-1.14	12 905 398	24 875 428	285 736 525	1 130 239 229	830 775 415	917 180 815	873 978 115	6,56
RF-1.16	13 120 569	25 603 284	295 445 530	1 125 240 154	824 654 406	915 117 733	869 886 070	6,63
RF-1.18	13 767 531	28 607 555	328 195 979	1 105 476 297	715 928 152	906 857 046	811 392 599	6,75
RF-1.20	13 816 305	28 717 283	329 993 753	1 105 107 596	722 534 022	906 726 593	814 630 308	6,76

Figure 137 Pit A4 Best-Worst Case Analysis

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Figure 138 A4 Pit Shell Strip Ratio and Grade

It is not immediately obvious from the Best-Worst Case analysis which pit shell(s) best represent the optimum final pit and which represent the best pushbacks to maximise NPV. This prompted some further analysis and to investigate some "specified cases" by selecting 2-3 interim pushbacks and final pit. The following specified cases sequence were selected for further analysis:

Specified Case 1: RF=0,60 – RF=0,80 – RF=0,98
Specified Case 2: RF=0,60 – RF=0,80 – RF=1,00
Specified Case 2b: RF=0,60 – RF=0,78 – RF=0,98
Specified Case 3: RF=0,60 – RF=0,84 – RF=0,98
Specified Case 3b: RF=0,60 – RF=0,84 – RF=1,00

The results from the 5 specified cases are tabulated below in and graphically illustrated in Figure 139. The results indicate that Specified Case 3 with RF=0,98 pit shell yields the best NPV and that RF=0,60 and RF=0,84 pit shell yields are the best interim pushbacks.

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Table 84 Specified Case Analysis

Specified Case 1
Revenue Factor	Pit Shell Inventory Mt	Incremental Pit Shell Mt	Net Value USD M	NPV_{R} USD M
0,60	69,0	69,0	987,9	832,4
0,80	122,1	53,1	1 085,5	893,4
0,98	232,6	110,5	1 136,8	905,1
Specified Case 2
Revenue Factor	Inventory Mt	Incremental Mt	Net Value USD M	NPV_{R} USD M
0,60	69,0	69,0	987,9	832,4
0,80	122,1	53,1	1 085,5	893,4
1,00	239,8	117,7	1 137,8	903,5
Specified Case 2b
Revenue Factor	Inventory Mt	Incremental Mt	Net Value USD M	NPV_{R} USD M
0,60	69,0	69,0	990,7	834,8
0,78	109,8	40,8	1 076,2	889,7
0,98	232,6	122,8	1 143,4	906,3
Specified Case 3
Revenue Factor	Inventory Mt	Incremental Mt	Net Value USD M	NPV_{R} USD M
0,60	69,0	69,0	990,7	834,8
0,84	132,2	63,2	1 101,1	899,0
0,98	232,6	100,4	1 143,4	908,5
Specified Case 3b
Revenue Factor	Inventory Mt	Incremental Mt	Net Value USD M	NPV_{R} USD M
0,60	69,0	69,0	990,7	834,8
0,84	132,2	63,2	1 101,1	899,0
1,00	239,8	107,6	1 144,5	905,9

Figure 139 Specified Case Analysis

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16.5.3. Final Pit and Interim Shells Selection

16.5.3.1. A2B Pit Selection

The optimum pit shell for pit A2B was selected to be the pit with a revenue factor of 0,92 (pit number 17). No interim pushbacks are selected. The selected pit shell contains 1,96 Mt of ore (Proven and Probable) at a TREO grade of 0,171% for 3 353 tonnes of total contained rare earth oxides. The strip ratio for the pit shell 8,5:1. See Table 85 for details.

Table 85 A2B Pit Shell RF=0,92 Pit Inventory
| Item | Unit | Value |
| --- | --- | --- |
| High-Grade | (t) | 940 235 |
| Low-Grade | (t) | 1 020 593 |
| Total Rom | (t) | 1 960 827 |
| Marginal | (t) | 517 041 |
| Waste | (t) | 16 111 888 |
| Total | (t) | 18 589 756 |
| Strip Ratio | Wst:Ore | 8,48 |
| High-Grade LREO | (t) | 1 000,1 |
| High-Grade HREO | (t) | 1 056,5 |
| High-Grade TREO | (t) | 2 056,6 |
| Low-Grade LREO | (t) | 498,6 |
| Low-Grade HREO | (t) | 797,7 |
| Low-Grade TREO | (t) | 1 296,3 |
| ROM LREO | (t) | 1 498,7 |
| ROM HREO | (t) | 1 854,1 |
| ROM TREO | (t) | 3 352,8 |
| High-Grade LREO | (%) | 0,106 |
| High-Grade HREO | (%) | 0,112 |
| High-Grade TREO | (%) | 0,219 |
| Low-Grade LREO | (%) | 0,049 |
| Low-Grade HREO | (%) | 0,078 |
| Low-Grade TREO | (%) | 0,127 |
| ROM LREO | (%) | 0,076 |
| ROM HREO | (%) | 0,095 |
| ROM TREO | (%) | 0,171 |

16.5.3.2. A4 Pit Selection

The optimum pit shell for pit A4 was selected to be the pit with a revenue factor of 0,98 (pit number 20). The selected pit shell contains 31,58 Mt of ore (Proven and Probable) at a TREO grade of 0,175% for 3 353 tonnes of total contained rare earth oxides. The strip ratio for the pit shell 6,4:1. See Table 86 for details.

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Table 86 A4 Pit Shell RF=0,98 Pit Inventory

Item	Unit	Value
High-Grade	(t)	11 266 971
Low-Grade	(t)	20 315 076
Total Rom	(t)	31 582 047
Marginal	(t)	10 773 598
Waste	(t)	190 303 502
Total	(t)	232 659 147
Strip Ratio	Wst:Ore	6,37
High-Grade LREO	(t)	10 015,1
High-Grade HREO	(t)	19 760,8
High-Grade TREO	(t)	29 775,9
Low-Grade LREO	(t)	13 669,0
Low-Grade HREO	(t)	11 711,1
Low-Grade TREO	(t)	25 379,8
ROM LREO	(t)	23 684,0
ROM HREO	(t)	31 471,9
ROM TREO	(t)	55 155,7
High-Grade LREO	(%)	0,089
High-Grade HREO	(%)	0,175
High-Grade TREO	(%)	0,264
Low-Grade LREO	(%)	0,067
Low-Grade HREO	(%)	0,058
Low-Grade TREO	(%)	0,125
ROM LREO	(%)	0,075
ROM HREO	(%)	0,100
ROM TREO	(%)	0,175

Figure 140 Final pit shells selected for A2B and A4 pits

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Two interim pushbacks for A4 pit were selected according to the specified cases analysed in Table 84. These two pit shells are RF=0,60 and RF=0,84. The pushbacks and final pit are illustrated in Figure 141 and the incremental pit shell inventory listed in Table 87. As expected, the increasing pushbacks increase in strip ratio and reduces in grade. It is obvious that 87% of the undiscounted value and 92% of the net present value is within the first pushback. The second and final pushback decision should be deferred as much as possible.

Figure 141 A4 Interim Pushbacks

Table 87 A4 Interim Pushbacks Summary

Item	Unit	RF=0,60	RF=0,84	RF=0,98	Total
High-Grade	(t)	6 562 918	2 117 743	2 586 310	11 266 971
Low-Grade	(t)	6 695 745	6 205 029	7 414 301	20 315 076
Total Rom	(t)	13 258 663	8 322 773	10 000 612	31 582 047
Marginal	(t)	2 758 284	3 954 091	4 061 223	10 773 598
Waste	(t)	53 000 080	50 932 435	86 370 988	190 303 502
Total	(t)	69 017 026	63 209 299	100 432 823	232 659 147
Strip Ratio	Wst:Ore	4,21	6,59	9,04	6,37
High-Grade LREO	(t)	5 076,1	2 230,8	2 708,1	10 015,1
High-Grade HREO	(t)	15 231,5	2 069,1	2 460,2	19 760,8
High-Grade TREO	(t)	20 307,6	4 299,9	5 168,3	29 775,9
Low-Grade LREO	(t)	3 949,4	4 590,2	5 129,4	13 669,0
Low-Grade HREO	(t)	4 476,7	3 107,5	4 126,8	11 711,1
Low-Grade TREO	(t)	8 426,1	7 697,6	9 256,1	25 379,8
ROM LREO	(t)	9 025,6	6 821,0	7 837,5	23 684,0
ROM HREO	(t)	19 708,2	5 176,6	6 587,1	31 471,9
ROM TREO	(t)	28 733,7	11 997,5	14 424,5	55 155,7
High-Grade LREO	(%)	0,077	0,105	0,105	0,089
High-Grade HREO	(%)	0,232	0,098	0,095	0,175
High-Grade TREO	(%)	0,309	0,203	0,200	0,264
Low-Grade LREO	(%)	0,059	0,074	0,069	0,067

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Item	Unit	RF=0,60	RF=0,84	RF=0,98	Total
Low-Grade HREO	(%)	0,067	0,050	0,056	0,058
Low-Grade TREO	(%)	0,126	0,124	0,125	0,125
ROM LREO	(%)	0,068	0,082	0,078	0,075
ROM HREO	(%)	0,149	0,062	0,066	0,100
ROM TREO	(%)	0,217	0,144	0,144	0,175
Net Value	USDM	990,7	110,5	42,2	1 143,4
Net Present Value	USDM	834,8	64,2	9,5	908,5

16.6. Open Pit Design

As standard practice in mine design, the optimised pit shells are used as templates to guide the pit design process. The pit shells are generated by following the recommended slope angles. Other design criteria like ramp widths and gradients, switchback radii, minimum pushback width and minimum mining width are included and applied during the pit design process.

Overall difference between pit shell and design is generally aimed to be ≤ 5% and ore volume be as to be ≤ 3%.

16.6.1. Pit Ramps Design Criteria

The pits wall configurations are as per Table 66 and Table 67 and the geotechnical design sectors in Figure 129 in section 16.2.

Additional design criteria are described in Figure 142.

Figure 142 Pit Ramp Design Criteria

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16.6.2. Pit A2B Design

Final design for pit A2B was developed as a single phase incorporating the ramp system as shown in Figure 143.

The inventory difference between the pit design and the pit shell for A2B compares well as shown in Table 88. The total tonnage difference between the pit design is +6.5% (500kt) with the total ROM difference being -3,0% (-58kt). There is some opportunity to further reduce the difference by incorporating the ramp system in the eastern end of the pit in the pit optimisation and to increase the pit size/reduce strip ratio by converting the resources to indicated and/or measured north of the current pit shell.

Figure 143 Pit A2B Final Pit Design

QUOSHO		QUOSHO UNION	Projected Lofdal Heavy Rare Earths Project Area A2B and 4+ Project Area
Decent by:	P Christiano	Date: 10 December 2008	Title: Pit A2B Final Pit
Checked by:	W Wosman	Drawing Number: NCML_1000_Dr

Table 88 A2B Pit Design vs Optimum Pit Shell

Item	Unit	A2B PD Phs_1	PS RF=0,98	Var	%Var
High-Grade	(t)	894 486	940 235	-45 749	-4,9%
Low-Grade	(t)	1 007 917	1 020 593	-12 676	-1,2%
Total Rom	(t)	1 902 402	1 960 827	-58 425	-3,0%
Marginal	(t)	578 094	517 041	61 053	11,8%
Waste	(t)	16 609 740	16 111 888	497 853	3,1%
Total	(t)	19 090 237	18 589 756	500 480	2,7%
Strip Ratio	Wst:Ore	9,03	8,48	-9,57	6,5%
High-Grade LREO	(t)	946,1	1 000,1	-54,0	-5,4%
High-Grade HREO	(t)	997,1	1 056,5	-59,4	-5,6%
High-Grade TREO	(t)	1 943,1	2 056,6	-113,4	-5,5%
Low-Grade LREO	(t)	484,2	498,6	-14,4	-2,9%
Low-Grade HREO	(t)	795,7	797,7	-2,0	-0,2%
Low-Grade TREO	(t)	1 279,9	1 296,3	-16,3	-1,3%
ROM LREO	(t)	1 430,3	1 498,7	-68,4	-4,6%
ROM HREO	(t)	1 792,8	1 854,1	-61,4	-3,3%

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Item	Unit	A2B PD Phs_1	PS RF=0,98	Var	%Var
ROM TREO	(t)	3 223,1	3 352,8	-129,7	-3,9%
High-Grade LREO	(%)	0,106	0,106	-0,001	-0,6%
High-Grade HREO	(%)	0,111	0,112	-0,001	-0,8%
High-Grade TREO	(%)	0,217	0,219	-0,001	-0,7%
Low-Grade LREO	(%)	0,048	0,049	-0,001	-1,7%
Low-Grade HREO	(%)	0,079	0,078	0,001	1,0%
Low-Grade TREO	(%)	0,127	0,127	-0,000	-0,0%
ROM LREO	(%)	0,075	0,076	-0,001	-1,6%
ROM HREO	(%)	0,094	0,095	-0,000	-0,3%
ROM TREO	(%)	0,169	0,171	-0,002	-0,9%

16.6.3. Pit A4 Design

Final design for pit A4 was developed as three phases incorporating the ramp systems as shown in Figure 144.

Figure 144 Pit A4 Interim Pushbacks and Final Pit Designs

The inventory difference between the Phase 3 (final) pit design and the pit shell for A4 compares well as shown in Table 89. The total tonnage difference between the pit design and selected pit shell is -1,5% (-3,6 Mt) with the total ROM difference being -4,8% (-1,5 Mt).

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Table 89 A4 Pit Design vs Optimum Pit Shell

Item	Unit	A4 PD Phs_3	PS RF=0,98	Var	%Var
High-Grade	(t)	10 982 308	11 266 971	-284 664	-2,5%
Low-Grade	(t)	19 098 945	20 315 076	-1 216 131	-6,0%
Total Rom	(t)	30 081 253	31 582 047	-1 500 794	-4,8%
Marginal	(t)	10 326 424	10 773 598	-447 174	-4,2%
Waste	(t)	188 686 174	190 303 502	-1 617 328	-0,8%
Total	(t)	229 093 850	232 659 147	-3 565 297	-1,5%
Strip Ratio	Wst:Ore	6,62	6,37	1,38	3,9%
High-Grade LREO	(t)	9 675,5	10 015,1	-339,5	-3,4%
High-Grade HREO	(t)	19 536,0	19 760,8	-224,8	-1,1%
High-Grade TREO	(t)	29 211,5	29 775,9	-564,3	-1,9%
Low-Grade LREO	(t)	12 769,7	13 669,0	-899,3	-6,6%
Low-Grade HREO	(t)	11 107,5	11 711,1	-603,6	-5,2%
Low-Grade TREO	(t)	23 877,0	25 379,8	-1 502,8	-5,9%
ROM LREO	(t)	22 445,2	23 684,0	-1 238,8	-5,2%
ROM HREO	(t)	30 643,5	31 471,9	-828,4	-2,6%
ROM TREO	(t)	53 088,5	55 155,7	-2 067,2	-3,7%
High-Grade LREO	(%)	0,088	0,089	-0,001	-0,9%
High-Grade HREO	(%)	0,178	0,175	0,002	1,4%
High-Grade TREO	(%)	0,266	0,264	0,002	0,6%
Low-Grade LREO	(%)	0,067	0,067	-0,000	-0,6%
Low-Grade HREO	(%)	0,058	0,058	0,001	0,9%
Low-Grade TREO	(%)	0,125	0,125	0,000	0,1%
ROM LREO	(%)	0,075	0,075	-0,000	-0,5%
ROM HREO	(%)	0,102	0,100	0,002	2,2%
ROM TREO	(%)	0,176	0,175	0,002	1,1%

Table 90 Pit A4 Design Inventory Summary

Item	Unit	A4 PD Phs_1	A4 PD Phs_2	A4 PD Phs_3	A4 PD_Phs1-3
High-Grade	(t)	4 999 893	2 924 067	3 059 703	10 983 663
Low-Grade	(t)	4 282 498	6 338 774	8 499 328	19 120 600
Total Rom	(t)	9 282 391	9 262 841	11 559 031	30 104 263
Marginal	(t)	1 907 505	3 428 849	4 993 623	10 329 977
Waste	(t)	47 896 726	55 475 303	85 353 954	188 725 983
Total	(t)	59 086 621	68 166 993	101 906 609	229 160 223
Strip Ratio	Wst:Ore	5,37	6,36	7,82	6,61
High-Grade LREO	(t)	3 653,9	2 921,8	3 101,3	9 676,9
High-Grade HREO	(t)	11 972,0	4 532,4	3 032,7	19 537,1
High-Grade TREO	(t)	15 625,8	7 454,2	6 133,9	29 213,9
Low-Grade LREO	(t)	2 409,1	4 485,4	5 891,9	12 786,4
Low-Grade HREO	(t)	3 005,0	3 426,6	4 686,1	11 117,7
Low-Grade TREO	(t)	5 414,0	7 911,9	10 578,0	23 903,9
ROM LREO	(t)	6 062,9	7 407,2	8 993,2	22 463,3
ROM HREO	(t)	14 976,9	7 959,0	7 718,8	30 654,8
ROM TREO	(t)	21 039,8	15 366,1	16 712,0	53 117,9
High-Grade LREO	(%)	0,073	0,100	0,101	0,088
High-Grade HREO	(%)	0,239	0,155	0,099	0,178
High-Grade TREO	(%)	0,313	0,255	0,200	0,266

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Low-Grade LREO	(%)	0,056	0,071	0,069	0,067
Low-Grade HREO	(%)	0,070	0,054	0,055	0,058
Low-Grade TREO	(%)	0,126	0,125	0,124	0,125
ROM LREO	(%)	0,065	0,080	0,078	0,075
ROM HREO	(%)	0,161	0,086	0,067	0,102
ROM TREO	(%)	0,227	0,166	0,145	0,176

16.7. Dump Design

The general waste rock dumps and stockpile design parameters are shown in Figure 145 below. The berm width and lift height allows for the final dump batters to be doxed at $20^{\circ}$ angle if required.

Figure 145 Dump Design Parameters

Figure 146 below shows the waste dumps and stockpiles around the site and how these are connected to the various pits and destinations.

Table 91 gives the capacities of the waste dumps and stockpiles.

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Figure 146 Waste Rock Dumps & Stockpiles
Table 91 Waste Dumps and Stockpiles Capacity

Dump/Stockpile	Lift Elevation	Capacity Volume m3	Capacity Tonnes
A2B Waste Dump	950	282 708	565 416
	970	5 187 578	10 375 156
	990	5 364 512	10 729 024
TOTAL A2B Waste Dump		10 834 798	21 669 596
A4 North Waste Dump	990	29 363 182	58 726 364
	1 010	14 207 527	28 415 054
	1 030	10 128 909	20 257 818
	1 050	6 548 472	13 096 944
Total A4 North Waste Dump		60 248 090	120 496 180
A4 South Waste Dump	970	620 702	1 241 404
	990	10 165 559	20 331 118
	1 010	13 288 897	26 577 794
	1 030	10 246 221	20 492 442
Total A4 South Waste Dump		34 321 379	68 642 758
A2B Marginal Stockpile	985	914 434	1 828 868
A4 Marginal Stockpile	1 000	5 351 638	10 703 276
L-G Fines Stockpile	1 020	1 998 201	3 996 402
A2B Topsoil Stockpile	960	52 733	84 373
A4 Topsoil Stockpile	975	202 993	324 789

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16.8. Dewatering

Due to the arid nature of the region, no serious dewatering issues are expected. However, minimal pumping capacity has been allowed for. Water ingress due to rainfall will be managed with berms and cut-off drains.

16.9. Operating Hours

Total yearly hours used was 8 760, with 730 hours per month. Weather consideration was calculated as follows:

Table 92 Weather Consideration

Weather
Rain factor	0,50	h/1mm of rain
Average rainfall/year	150	mm/year
Hours lost/year	75	h/year
Hours/day	24	h
Days lost/year	3,13	days/ year

Three public holidays (Christmas Day, Family Day and New Year's Day) were added as stoppage time.

Other factors that were included were mobilise and start-up checks, lunch and other breaks, end of shift, blasting and demobilising time. Adding Utilisation and Availability together with overall job efficiency results in 476 actual working hours per month. This was then used in the equipment calculations.

16.10. Mining Equipment

Lofdal will be mined as a conventional shovel and truck operation, with bulk mining augmented by more selective mining in areas with narrow ore zones and ore/waste contact zones. It was assumed whole mining operation will be outsourced to a reputable mining contractor Service Provider (Contractor), except for the mine technical services function. This includes drilling, blasting, loading and hauling of ore and waste. The merits of contract mining include:

Contractors can quickly deploy equipment and specialist work force addressing skills, staffing and equipment shortages.
Junior mining companies could increase their scale of operations without making large investments in capital and labour.
Under-capitalized mining companies are provided with the means to develop their mines more rapidly and cheaply than if they had relied on conventional sources of finance.
Reputable contract mining companies provide mines with improved operational and best practice, cost efficiency, and effective performance and safety management systems.

The condensed preferred mobile mining equipment is shown in Table 93.

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Table 93 Condensed Mining Equipment List
| Equipment | Pre-strip | Year 1 | Year 2 | Year 3 | Year 4 | Year 5 |
| --- | --- | --- | --- | --- | --- | --- |
| Production Excavator - option 1 | 2 | 4 | 4 | 4 | 4 | 4 |
| Production Excavator - option 2 | 1 | 1 | 1 | 1 | 1 | 1 |
| Front-end Wheel Loaders | 1 | 2 | 2 | 2 | 2 | 2 |
| Haul Trucks | 8 | 18 | 19 | 20 | 21 | 22 |
| Drill Rigs | 3 | 4 | 5 | 6 | 6 | 6 |
| Secondary Excavators | 2 | 2 | 2 | 2 | 2 | 2 |
| Track Dozers | 3 | 4 | 4 | 4 | 4 | 4 |
| Wheel Dozers | 0 | 0 | 0 | 0 | 0 | 0 |
| Graders | 1 | 1 | 2 | 2 | 2 | 2 |
| Diesel Bowser | 1 | 1 | 1 | 1 | 1 | 1 |
| Water Carts | 2 | 2 | 2 | 2 | 2 | 2 |
| LDV & Busses | 26 | 28 | 28 | 30 | 30 | 30 |
| Maintenance Support Equipment | 6 | 6 | 6 | 6 | 6 | 6 |
| Blasting Support Equipment | 8 | 10 | 10 | 10 | 10 | 10 |
| Road Maintenance | 2 | 2 | 2 | 2 | 2 | 2 |
| Other | 8 | 10 | 11 | 11 | 12 | 12 |
| Total Item A2 | 74 | 95 | 99 | 103 | 105 | 106 |

16.11. Mining Personnel

16.11.1.1. Owner's Mining Labour Complement

The project owner provides all technical services to direct and support the mining contractor. The owner's mining personnel complement is detailed in Table 94 below. Total labour complement for technical services is expected to stabilise at 23 pax at steady state production.

Table 94 Owner's Mining Labour Complement
| Mining Owner Team Numbers | Unit | PS Q2 | PS Q1 | Y1 Q1 | Y1 Q2 | Y1 Q3 | Y1 Q4 | Y2 Q5 | Y2 Q6 | Y2 Q7 | Y2 Q8 | Y3 Q9 | Y3 Q10 | Y3 Q11 | Y4 Q12 | Y5 | Y6 | Y7 | Y8 | Y9 | Y10 | Y11 | Y12 | Y13 |
| --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- | --- |
| Mining Manager | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Supt Tech Services (Chief Geo) | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Grade Control Geologist | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Production Control Geologist | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Geology Field Assistant | nr | 2 | 2 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 |
| Resource Geologist | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Geotechnical Engineer | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Geotechnical Field Assistant | nr | 1 | 1 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 |
| Senior Mine Surveyor | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Mine Surveyor | nr | 1 | 1 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 |
| Surveyor Assistant | nr | 2 | 2 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 |
| Draftsman | nr | 1 | 1 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 | 3 |
| Supt Mining (Chief Mining Eng) | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Short Term Planning Engineer | nr | 1 | 1 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 | 2 |
| Medium/Long Term Planning Eng | nr | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 | 1 |
| Total | | 16 | 16 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 | 23 |

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16.11.1.2. Mining Contractor Labour Complement

The mining contractor labour complement was calculated using a 3-shift FULCO system. There are 3 operators allowed for each piece of production equipment excluding annual, sick leave and absence. The next Table 95 to Table 98 gives the detail of the labour complement required by the contractor for the Management and Administration-, Production- and Maintenance Staff. Total contractor labour is expected to start at 326 personnel during the pre-strip period and increase to full complement of 574 personnel by production year-5.

Table 95 Mining Contractor Management and Admin Labour Complement

B1.1	Management & Administration Staff
	Job Description / Job Title		Paterson Scale	Unit	Pre-Strip	Y01	Y02	Y03	Y04	Y05
B1.1.1	Contracts Manager	Operational GM	E3	nr	1	1	1	1	1	1
B1.1.2	Planning Engineer	Snr Professional/Supt.	D2	nr	1	1	1	1	1	1
B1.1.3	Drilling Engineer	Snr Professional/Supt.	D2	nr	1	1	1	1	1	1
B1.1.4	Drilling Foreman	Foreman/Graduate	C4	nr	4	4	4	4	4	4
B1.1.5	Drilling Supervisor	Supervisor/Artisan	C2	nr	0	0	0	0	0	0
B1.1.6	Blasting/Explosive Engineer	Snr Professional/Supt.	D2	nr	1	1	1	1	1	1
B1.1.7	Blasting Foreman	Foreman/Graduate	C4	nr	2	2	2	2	2	2
B1.1.8	Blasting Supervisor	Supervisor/Artisan	C2	nr	0	0	0	0	0	0
B1.1.9	L&H Engineer	Snr Professional/Supt.	D2	nr	1	1	1	1	1	1
B1.1.10	L&H Foreman	Foreman/Graduate	C4	nr	1	1	1	1	1	1
B1.1.11	L&H Supervisor	Supervisor/Artisan	C2	nr	4	8	8	8	8	8
B1.1.12	Surveyor	Professional	D1	nr	2	3	3	3	3	3
B1.1.13	Surveyor Assistant	Unskilled <5y experience	A2	nr	0	0	0	0	0	0
B1.1.14	Dispatcher	Junior Professional	C5	nr	1	1	1	1	1	1
B1.1.15	Financial/ Commercial Manager	Technical Manager	D4	nr	0	0	0	0	0	0
B1.1.16	Site Accountant	Professional	D1	nr	1	1	1	1	1	1
B1.1.17	Commercial Officer	Professional	D1	nr	0	0	0	0	0	0
B1.1.18	HR Manager	Technical Manager	D4	nr	0	0	0	0	0	0
B1.1.19	HR Officer	Professional	D1	nr	2	3	3	3	3	3
B1.1.20	SHE Manager	Technical Manager	D4	nr	1	1	1	1	1	1
B1.1.21	SHE Officer	Professional	D1	nr	1	2	2	2	2	2
B1.1.22	Office & General	Junior Professional	C5	nr	1	1	1	1	1	1
B1.1.23	Training Instructor	Snr Professional/Supt.	D2	nr	1	1	1	1	1	1
B1.1.24	Office and Change house Cleaners	Unskilled - no experience	A1	nr	8	12	12	12	12	12
B1.1.25	Change house Laundry Attendants	Semi-Skilled <5y experience	B2	nr	0	0	0	0	0	0
B1.1.26	Clerks/Pit Controller	Semi-Skilled <5y experience	B2	nr	10	12	12	12	12	12
	Sub-total Item B1.1			nr	44	57	57	57	57	57

Table 96 Mining Contractor Production Staff Labour Complement

B1.2	Production Staff
	Job Description / Job Title		Paterson Scale	Unit	Pre-Strip	Y01	Y02	Y03	Y04	Y05
B1.2.1	Production Excavator Operator	Snr Operator/Blaster	B5	nr	16	28	32	32	32	32
B1.2.2	Front Loader Operator	Operator	B4	nr	4	12	12	12	12	12
B1.2.3	Haul Truck Operator	Operator	B4	nr	64	152	188	196	196	208
B1.2.4	Drill Rig Operator	Snr Operator/Blaster	B5	nr	14	28	28	32	32	32
B1.2.5	Excavator with hydraulic hammer (rock breaker) Operator	Operator	B4	nr	4	4	4	4	4	4
B1.2.6	Excavator for highwall scaling(long-reach) Operator	Operator	B4	nr	4	4	4	4	4	4
B1.2.7	Track Dozer Operator	Operator	B4	nr	20	24	28	28	28	32
B1.2.8	Wheel Dozer Operator	Operator	B4	nr	0	0	0	0	0	0
B1.2.9	Grader Operator	Operator	B4	nr	4	4	8	8	8	8
B1.2.10	Diesel Bowser Driver	Operator	B4	nr	4	4	4	4	4	4
B1.2.11	Water Carts Driver	Operator	B4	nr	8	8	8	8	8	8
B1.2.12	Road Sweeper truck Driver	Operator	B4	nr	0	0	0	0	0	0
B1.2.13	Lube truck/Flat-bed Driver	Operator	B4	nr	6	6	6	6	6	6
B1.2.14	Shift Change Bus/Shuttle/Quantum Driver	Operator	B4	nr	12	12	12	12	12	12
B1.2.15	Crusher operator	Operator	B4	nr	12	12	12	12	12	12

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B1.2	Production Staff
	Job Description / Job Title		Paterson Scale	Unit	Pre-Strip	Y01	Y02	Y03	Y04	Y05
B1.2.16	Aggregate truck driver	Operator	B4	nr	8	8	8	8	8	8
B1.2.17	Drill Assistant	Semi-Skilled - no experience	B1	nr	14	28	28	32	32	32
B1.2.18	Ore spotter	Semi-Skilled - no experience	B1	nr	12	12	12	12	12	12
B1.2.19	Cleaner	Semi-Skilled - no experience	B1	nr	8	8	8	8	8	8
B1.2.20	MMU Operator	Operator	B4	nr	4	4	4	4	4	4
B1.2.21	Blaster	Snr Operator/Blaster	B5	nr	4	4	4	4	4	4
B1.2.22	Blaster Assistants	Semi-Skilled - no experience	B1	nr	12	12	12	12	12	12
	Sub-total Item B1.2			nr	234	374	422	438	438	454

Table 97 Mining Contractor Maintenance Staff Labour Complement

B1.3	Maintenance Staff
	Job Description / Job Title		Paterson Scale	Unit	Pre-Strip	Y01	Y02	Y03	Y04	Y05
B1.3.1	Maintenance Manager	Technical GM	E2	nr	1	1	1	1	1	1
B1.3.2	Maintenance Engineer	Snr Professional/Supt.	D2	nr	0	0	0	0	0	0
B1.3.3	Workshop Foreman	Foreman/Graduate	C4	nr	1	1	2	2	2	2
B1.3.4	Artisans - Mechanical	Snr Artisan	C3	nr	7	7	8	10	11	11
B1.3.5	Artisans - Electrical	Snr Artisan	C3	nr	1	1	1	1	1	1
B1.3.6	Artisans - Boilermaker	Snr Artisan	C3	nr	2	2	2	2	2	2
B1.3.7	Other Senior Artisans	Snr Artisan	C3	nr	0	0	0	0	0	0
B1.3.8	Artisan Assistance	Semi-Skilled <5y experience	B2	nr	10	10	11	13	14	14
B1.3.9	Storeman	Semi-Skilled >5y experience	B3	nr	1	1	1	1	1	1
B1.3.10	Tyre Bay Supervisor	Supervisor/Artisan	C2	nr	1	1	1	1	1	1
B1.3.11	Tyre Bay & Wash Bay Assistance	Semi-Skilled <5y experience	B2	nr	4	4	4	4	4	4
B1.3.12	Plant Clerk	Junior Professional	C5	nr	1	1	1	1	1	1
B1.3.13	Drill Maintenance Supervisor	Professional	D1	nr	1	1	1	1	1	1
B1.3.14	Drill Maintenance Artisans	Supervisor/Artisan	C2	nr	5	7	7	8	8	8
B1.3.15	Drill Maintenance assistance	Semi-Skilled >5y experience	B3	nr	5	7	7	8	8	8
B1.3.16	GET Operative	Semi-Skilled >5y experience	B3	nr	4	4	4	4	4	4
B1.3.17	Admin Clerk	Semi-Skilled <5y experience	B2	nr	4	4	4	4	4	4
	Sub-total Item B1.3			nr	48	52	55	61	63	63

Table 98 Mining Contractor Total Labour Complement

B1	Mining Contractor Labour Complement
	Area	Unit	Pre-Strip	Y01	Y02	Y03	Y04	Y05
B1.1	Management & Administration Staff	nr	44	57	57	57	57	57
B1.2	Production Staff	nr	234	374	422	438	438	454
B1.3	Maintenance Staff	nr	48	52	55	61	63	63
B1.4	Total Item B1	nr	326	483	534	556	558	574

16.12. Production Schedule

The mining production schedule was accomplished with the MinePlan Schedule Optimiser (MPSO). Production periods for the were setup to do the pre-strip in two quarters followed by 12 quarterly increments for the first three production years, with the rest of the schedule done in annual increments.

16.12.1. Scheduler Setup

The pit designs were used to generate mining cuts within each pit phase. The mining cuts were generated for 10m benches along strike of the ore body and target cut size of 25 000 m³ (approximately 70 000 tonnes) per cut, resulting in 3 003 total mining cuts.

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The scheduler process flow is depicted in Figure 147. The process flow of ROM includes stockpiles for High-Grade and Low-Grade to smooth out mining rate. ROM High-Grade and Low-Grade is fed to the respective crushers with front-end loaders operated by the plant. The Low-Grade stream includes a Process Point which includes the fines screening (<10mm) prior to the >10mm fraction being sent to the Ore Sorter. A Python script was developed and tested to calculate the volumes, tonnes and grades the various Low-Grade streams.

Figure 147 Scheduler (MPSO) Process Flow

16.12.2. Schedule Control

The primary schedule control is that of preventing undercutting upper benches within the same pit or pushback. The other key control is to maintain a ≥2-bench lag in the outer pushback of two adjacent phases.

The material mapping is also important to allow only the correct material to be sent for processing and other materials are routed to the appropriate destinations. The material mapping for the scheduler was done on:

MZONE – Mineralised zone
GTYPE – Grade Type classification
WEATH – Weathering classification
CLASS – Resource classification
TREO% - TREO grade

The vertical advance rate in each pushback is also limited to ≤7-benches per year. This is based on experience considering the bench height, bench area, and the drill/blast, load/haul cycle turnover.

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16.12.3. Schedule Targets

The primary scheduling objectives were to achieve the projected ROM tonnage feed for both High-Grade and Low-Grade materials. The McNulty 2 ramp-up curve was applied to guide the feed rates for High-Grade and Low-Grade ores, as well as to forecast flotation recoveries for HREO and LREO. Nameplate processing rates and recoveries are anticipated to be reached within eight quarters (32 months) from commissioning, as illustrated in Figure 148.

Furthermore, the multi-period scheduling strategy focused on maximising net present value (NPV). The overall mining rate was systematically managed to support a steady ramp-up and to minimize fluctuations in operational rates.

Figure 148 Process Throughput and Recovery Ramp-up

16.12.4. Schedule Results

The schedule includes all pit material movement, stockpile rehandling and -reclaiming, High-Grade crusher feed, Low-Grade crusher feed, Low-Grade fines stockpiling and waste reject from the XRT ore-sorter to the waste rock dump.

Though the tabulated schedule results below do not all report the total suite of rare-earths metal grades and content, all these values, these values are always tracked in the schedule and maybe reported when required.

The final vertical advance rates achieved in the schedule exceeds the targeted rates of seven benches per year for some of the pushbacks in some periods as shown in Table 99 below.

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Whilst not as a significant concern, this is noted and will be addressed with more detailed scheduling and cut sequencing.

Table 99 Vertical Advance Rates Achieved

Pit Phase	Number of Benches Mined
Mine Period	PS	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13
A2B Phs1						3	3	3	4	2	5
A4 Phs1	5	5	4	6	6	6	7
A4 Phs2		4	3	3	6	6	6	8	8	6	3
A4 Phs4				4	5	6	5	6	3	6	7	10	11	9

The table below summarise the main results from the mine production scheduling. These are shown in the next Table 101 through Table 110 and Figure 149 through Figure 151.

Table 100 Summary of the Results Tables and Figures

Table / Figure	Comments
Table 101 Pre-Strip and First 3-Years Quarterly Pit Production Schedule	Mined tonnage and grades for the pre-strip period and first three years of production with annual sub-totals.
Figure 149 Pre-strip and First 3-Years Quarterly Pit Production Schedule	Same as above presented graphically
Table 102 LOM Annual Pit Production Schedule	Mined tonnage and grades for the pre-strip period and production periods presented annually
Figure 150 LOM Annual Pit Production Schedule	Same as above presented graphically
Table 103 Pre-Strip and First 3-Years Quarterly Ore Mined by Measured and Indicated Resource Class	Pre-strip period and first three years ore tonnes mined, and grade presented in measured and indicated resource classification
Table 104 LOM Annual Ore Mined by Measured and Indicated Resource Class	LOM annual ore tonnes mined, and grade presented in measured and indicated resource classification
Figure 151 ROM Mined by Resource Class	Same as the two items above presented graphically
Table 105 LOM Ex-Pit ROM to Destinations	LOM ROM tonnage mined and grades by destination (Process Plant, Ore-Sorting Plant, High- and Low-Grade stockpiles). In Years 7 to -11, and Year 12, some 765kt Low-Grade material is fed to Mill/Flotation plant as there was not enough High-Grade material available.
Table 106 LOM Stockpile Reclaim to Mill/Flotation	LOM ore material reclaimed from the ore stockpiles plus concentrate from the XRT plant all feed to the Mill/Flotation plant
Table 107 LOM Marginal Stockpiled	LOM Marginal stockpiled from the 2 mining pits
Table 108 LOM XRT Reject to A4N Waste Rock Dump	LOM XRT reject tonnage and grades sent to the A4-North waste dump
Table 109 LOM Waste to Waste Rock Dumps	LOM tonnage transported to the various waste rock dumps
Table 110 LOM Annual Mill/Flotation Feed	LOM detail source and grades of Mill/Flotation feed for metal recovery.

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Table 101 Pre-Strip and First 3-Years Quarterly Pit Production Schedule

			Pre-strip		Pre-Strip Total	Y01				Y01 Total	Y02				Y02 Total	Y03				Y03 Total	Grand Total
Pit	Material	Unit	PS-Q1	PS-Q2		Y01-Q1	Y01-Q2	Y01-Q3	Y01-Q4		Y02-Q1	Y02-Q2	Y02-Q3	Y02-Q4		Y03-Q1	Y03-Q2	Y03-Q3	Y03-Q4
Pit_A4	High Grade	(t) Tonnes	95 317	246 242	341 560	98 295	89 952	214 598	278 495	681 340	147 205	57 495	223 958	368 649	797 308	313 165	316 001	299 226	275 000	1 203 392	3 023 600
		(%) LREO	0,076	0,072	0,073	0,069	0,060	0,071	0,077	0,072	0,075	0,089	0,077	0,098	0,088	0,088	0,089	0,075	0,068	0,080	0,079
		(%) HREO	0,202	0,226	0,219	0,159	0,198	0,203	0,292	0,232	0,150	0,132	0,155	0,183	0,165	0,211	0,214	0,222	0,209	0,214	0,206
		(%) TREO	0,278	0,298	0,292	0,228	0,258	0,273	0,369	0,304	0,225	0,220	0,233	0,281	0,253	0,299	0,303	0,297	0,277	0,295	0,286
	Low Grade	(t) Tonnes	68 524	190 642	259 167	141 555	78 831	232 094	322 080	774 559	160 783	230 174	307 015	383 146	1 081 119	318 528	318 006	372 621	382 000	1 391 155	3 506 000
		(%) LREO	0,060	0,059	0,060	0,077	0,062	0,056	0,060	0,062	0,071	0,075	0,056	0,069	0,067	0,057	0,063	0,062	0,065	0,062	0,063
		(%) HREO	0,070	0,068	0,068	0,050	0,065	0,070	0,065	0,064	0,053	0,047	0,070	0,056	0,057	0,071	0,064	0,063	0,061	0,065	0,062
		(%) TREO	0,130	0,127	0,128	0,127	0,127	0,126	0,125	0,126	0,124	0,122	0,126	0,125	0,124	0,128	0,126	0,125	0,126	0,126	0,126
	Marginal	(t) Tonnes	32 126	96 890	129 017	70 061	30 018	124 471	174 259	398 809	87 660	150 318	125 099	242 575	605 651	131 997	144 289	203 910	246 859	727 055	1 860 532
		(%) LREO	0,049	0,043	0,044	0,052	0,046	0,043	0,048	0,047	0,056	0,057	0,045	0,058	0,055	0,047	0,050	0,053	0,053	0,052	0,051
		(%) HREO	0,041	0,047	0,045	0,038	0,043	0,047	0,042	0,043	0,035	0,033	0,045	0,032	0,035	0,044	0,041	0,036	0,036	0,038	0,039
		(%) TREO	0,090	0,089	0,089	0,091	0,089	0,089	0,090	0,090	0,091	0,090	0,090	0,090	0,090	0,091	0,090	0,090	0,089	0,090	0,090
	Waste	(t) Tonnes	1 604 032	2 466 225	4 070 257	2 690 089	2 801 198	2 628 838	2 725 166	10 845 291	3 904 351	4 162 013	4 243 927	4 105 631	16 415 922	4 536 310	4 721 704	4 799 718	4 765 599	18 823 331	50 154 802
Pit_A4 Total	(t) Tonnes	1 800 000	3 000 000	4 800 000	3 000 000	3 000 000	3 200 000	3 500 000	12 700 000	4 300 000	4 600 000	4 900 000	5 100 000	18 900 000	5 300 000	5 500 000	5 675 475	5 669 458	22 144 934	58 544 934
		Strip Ratio	9,986	5,87	6,99	11,51	16,77	6,16	4,83	7,72	12,96	14,99	8,23	5,78	9,06	7,39	7,67	7,45	7,63	7,54	7,97

Figure 149 Pre-strip and First 3-Years Quarterly Pit Production Schedule

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Table 102 LOM Annual Pit Production Schedule

PIT	Material	Unit	PS	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Grand Total
Pit_A2B	High Grade	(t) Tonnes						35 286	112 594	186 917	255 137	60 369	243 921	258			894 481
(%) LREO						0,111	0,107	0,104	0,089	0,092	0,126	0,171			0,106
(%) HREO						0,112	0,104	0,114	0,121	0,119	0,101	0,098			0,111
(%) TREO						0,223	0,211	0,218	0,210	0,211	0,227	0,269			0,217
Low Grade	(t) Tonnes						39 622	109 517	183 731	327 291	87 113	260 514	130			1 007 918
(%) LREO						0,053	0,050	0,050	0,045	0,041	0,051	0,066			0,048
(%) HREO						0,074	0,077	0,078	0,081	0,085	0,077	0,066			0,079
(%) TREO						0,127	0,127	0,128	0,126	0,126	0,128	0,132			0,127
Marginal	(t) Tonnes						22 335	67 131	110 373	169 155	54 026	155 003	70			578 094
(%) LREO						0,033	0,039	0,035	0,034	0,030	0,034	0,045			0,034
(%) HREO						0,059	0,051	0,055	0,055	0,060	0,056	0,045			0,055
(%) TREO						0,092	0,089	0,090	0,089	0,090	0,089	0,090			0,090
Min Waste	(t) Tonnes						41 886	85 541	131 023	218 157	69 083	198 215	88			743 993
Waste	(t) Tonnes						501 604	3 820 112	3 845 064	5 081 194	897 123	1 647 397	683			15 793 177
Pit_A2B Total	(t) Tonnes						640 733	4 194 895	4 457 108	6 050 935	1 167 714	2 505 050	1 229			19 017 664
Strip Ratio						7,55	17,89	11,03	9,39	6,92	3,97	2,17	0,00	0,00	9,00
Pit_A4	High Grade	(t) Tonnes	341 560	681 340	797 308	1 203 392	1 100 000	846 514	1 163 703	404 986	826 535	947 623	583 339	540 938	1 139 446	406 976	10 983 660
(%) LREO	0,073	0,072	0,088	0,080	0,082	0,087	0,082	0,108	0,092	0,098	0,083	0,114	0,088	0,111	0,088
(%) HREO	0,219	0,232	0,165	0,214	0,200	0,173	0,257	0,094	0,144	0,199	0,173	0,090	0,116	0,093	0,178
(%) TREO	0,292	0,304	0,253	0,295	0,282	0,260	0,339	0,202	0,237	0,297	0,256	0,204	0,204	0,204	0,266
Low Grade	(t) Tonnes	259 167	774 559	1 081 119	1 391 155	1 528 000	1 554 442	1 579 259	1 042 897	1 384 742	1 768 121	2 149 277	1 400 200	2 592 622	615 014	19 120 573
(%) LREO	0,060	0,062	0,067	0,062	0,071	0,069	0,066	0,073	0,064	0,070	0,070	0,066	0,062	0,073	0,067
(%) HREO	0,068	0,064	0,057	0,065	0,055	0,054	0,059	0,052	0,062	0,054	0,053	0,058	0,064	0,057	0,058
(%) TREO	0,128	0,126	0,124	0,126	0,125	0,123	0,125	0,125	0,126	0,124	0,123	0,124	0,126	0,130	0,125
Marginal	(t) Tonnes	129 017	398 809	605 651	727 055	1 062 394	947 673	636 575	607 058	694 881	1 059 958	1 378 018	865 692	1 084 828	129 260	10 326 868
(%) LREO	0,044	0,047	0,055	0,052	0,053	0,051	0,049	0,052	0,049	0,050	0,051	0,050	0,047	0,053	0,050
(%) HREO	0,045	0,043	0,035	0,038	0,036	0,039	0,041	0,037	0,041	0,040	0,039	0,040	0,043	0,037	0,039
(%) TREO	0,089	0,090	0,090	0,090	0,089	0,090	0,090	0,090	0,090	0,090	0,090	0,090	0,090	0,091	0,090
Min Waste	(t) Tonnes
Waste	(t) Tonnes	4 070 257	10 845 291	16 415 922	18 823 331	20 032 606	19 533 638	15 948 569	16 810 951	14 537 908	18 879 584	13 384 316	12 691 941	6 518 619	212 052	188 704 986
Pit_A4 Total	(t) Tonnes	4 800 000	12 700 000	18 900 000	22 144 934	23 723 000	22 882 267	19 328 105	18 865 892	17 444 065	22 655 286	17 494 950	15 498 771	11 335 515	1 363 302	229 136 087
Strip Ratio	6,99	7,72	9,06	7,54	8,03	8,53	6,05	12,03	6,89	7,34	5,40	6,98	2,04	0,33	6,61

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PIT	Material	Unit	PS	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Grand Total
Total Lofdal	High Grade	(t) Tonnes	5	681 340	797 308	1 203 392	1 100 000	881 800	1 276 297	591 903	1 081 672	1 007 992	827 259	541 195	1 139 446	406 976	11 878 141
		(%) LREO	0,073	0,072	0,088	0,080	0,082	0,088	0,084	0,107	0,092	0,098	0,096	0,114	0,088	0,111	0,089
		(%) HREO	0,219	0,232	0,165	0,214	0,200	0,170	0,244	0,100	0,139	0,194	0,152	0,090	0,116	0,093	0,173
		(%) TREO	0,292	0,304	0,253	0,295	0,282	0,258	0,328	0,207	0,231	0,292	0,248	0,204	0,204	0,204	0,262
	Low Grade	(t) Tonnes	259 167	774 559	1 081 119	1 391 155	1 528 000	1 594 064	1 688 775	1 226 628	1 712 033	1 855 234	2 409 791	1 400 330	2 592 622	615 014	20 128 491
		(%) LREO	0,060	0,062	0,067	0,062	0,071	0,069	0,065	0,069	0,061	0,069	0,068	0,066	0,062	0,073	0,066
		(%) HREO	0,068	0,064	0,057	0,065	0,055	0,055	0,060	0,056	0,066	0,055	0,055	0,058	0,064	0,057	0,059
		(%) TREO	0,128	0,126	0,124	0,126	0,125	0,123	0,125	0,125	0,126	0,124	0,124	0,124	0,126	0,130	0,125
	Marginal	(t) Tonnes	129 017	398 809	605 651	727 055	1 062 394	970 008	703 706	717 431	864 036	1 113 984	1 533 021	865 762	1 084 828	129 260	10 904 963
		(%) LREO	0,044	0,047	0,055	0,052	0,053	0,051	0,048	0,050	0,046	0,049	0,049	0,050	0,047	0,053	0,050
		(%) HREO	0,045	0,043	0,035	0,038	0,036	0,039	0,042	0,040	0,044	0,041	0,041	0,040	0,043	0,037	0,040
		(%) TREO	0,089	0,090	0,090	0,090	0,089	0,090	0,090	0,090	0,090	0,090	0,090	0,090	0,090	0,091	0,090
Min Waste Waste	(t) Tonnes						41 886	85 541	131 023	218 157	69 083	198 215	88			743 993
	(t) Tonnes	4 070 257	10 845 291	16 415 922	18 823 331	20 032 606	20 035 242	19 768 681	20 656 015	19 619 102	19 776 707	15 031 714	12 692 624	6 518 619	212 052	204 498 163
Total Lofdal	(t) Tonnes	4 800 000	12 700 000	18 900 000	22 144 934	23 723 000	23 523 000	23 523 000	23 323 000	23 495 000	23 823 000	20 000 000	15 500 000	11 335 515	1 363 302	248 153 751
		Strip Ratio	6,99	7,72	9,06	7,54	8,03	8,50	6,93	11,83	7,41	7,32	5,18	6,98	2,04	0,33	6,75

Figure 150 LOM Annual Pit Production Schedule

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Table 103 Pre-Strip and First 3-Years Quarterly Ore Mined by Measured and Indicated Resource Class

Resource Class	Material Type	Unit	PS Q1	PS Q2	Pre-Strip Total	Y01-Q1	Y01-Q2	Y01-Q3	Y01-Q4	Year-01 Total	Y02-Q1	Y02-Q2	Y02-Q3	Y02-Q4	Year-02 Total	Y03-Q1	Y03-Q2	Y03-Q3	Y03-Q4	Year-03 Total	Total
Measured	High Grade	Tonnes	78 310	235 287	313 598	76 143	89 772	207 270	238 304	611 489	130 388	25 524	195 655	261 777	613 343	286 156	282 329	249 161	169 149	986 796	2 525 225
(%) LREO	0,065	0,068	0,067	0,064	0,060	0,070	0,067	0,066	0,068	0,054	0,072	0,089	0,078	0,088	0,088	0,077	0,065	0,081	0,075
(%) HREO	0,224	0,233	0,231	0,177	0,198	0,205	0,329	0,249	0,161	0,190	0,162	0,222	0,189	0,221	0,227	0,233	0,228	0,227	0,223
(%) TREO	0,289	0,301	0,298	0,241	0,258	0,275	0,396	0,315	0,230	0,244	0,234	0,311	0,266	0,309	0,315	0,309	0,293	0,308	0,298
Low Grade	Tonnes	44 352	166 155	210 507	39 699	73 824	209 804	214 783	538 110	106 914	52 931	201 155	210 154	571 154	249 556	177 598	195 091	120 514	742 759	2 062 531
(%) LREO	0,055	0,054	0,055	0,062	0,060	0,055	0,053	0,056	0,061	0,065	0,054	0,063	0,060	0,057	0,054	0,052	0,050	0,054	0,056
(%) HREO	0,076	0,073	0,073	0,064	0,067	0,071	0,073	0,071	0,064	0,058	0,073	0,062	0,066	0,072	0,074	0,078	0,079	0,075	0,071
(%) TREO	0,131	0,127	0,128	0,126	0,127	0,126	0,126	0,126	0,126	0,123	0,126	0,125	0,126	0,129	0,128	0,130	0,129	0,129	0,127
Measured	Total	Tonnes	122 662	401 443	524 105	115 841	163 596	417 075	453 087	1 149 599	237 301	78 455	396 810	471 931	1 184 497	535 712	459 927	444 252	289 663	1 729 555	4 587 756
(%) LREO	0,062	0,062	0,062	0,063	0,060	0,063	0,060	0,061	0,065	0,062	0,063	0,077	0,069	0,074	0,074	0,066	0,058	0,069	0,066
(%) HREO	0,170	0,167	0,168	0,138	0,139	0,138	0,208	0,166	0,118	0,101	0,117	0,151	0,129	0,152	0,168	0,165	0,166	0,162	0,155
(%) TREO	0,232	0,229	0,230	0,202	0,199	0,200	0,268	0,227	0,183	0,163	0,179	0,228	0,198	0,225	0,242	0,230	0,225	0,231	0,221
Indicated	High Grade	Tonnes	17 007	10 955	27 962	22 152	180	7 328	40 191	69 852	16 818	31 971	28 303	106 872	183 964	27 009	33 673	50 065	105 851	216 597	498 375
(%) LREO	0,124	0,166	0,140	0,086	0,131	0,087	0,140	0,117	0,127	0,116	0,113	0,123	0,120	0,084	0,100	0,069	0,072	0,077	0,102
(%) HREO	0,104	0,072	0,091	0,098	0,035	0,130	0,075	0,088	0,060	0,085	0,108	0,088	0,088	0,112	0,105	0,169	0,178	0,157	0,118
(%) TREO	0,227	0,238	0,231	0,185	0,166	0,217	0,214	0,205	0,188	0,202	0,221	0,210	0,208	0,196	0,205	0,239	0,251	0,234	0,220
Low Grade	Tonnes	24 172	24 487	48 659	101 856	5 007	22 289	107 297	236 449	53 870	177 243	105 860	172 992	509 964	68 972	140 408	177 531	261 486	648 396	1 443 469
(%) LREO	0,069	0,093	0,081	0,083	0,095	0,058	0,075	0,077	0,091	0,078	0,062	0,077	0,076	0,059	0,074	0,073	0,071	0,071	0,074
(%) HREO	0,058	0,035	0,046	0,044	0,029	0,066	0,048	0,048	0,030	0,043	0,064	0,048	0,048	0,065	0,051	0,048	0,053	0,052	0,050
(%) TREO	0,127	0,128	0,128	0,127	0,123	0,124	0,123	0,125	0,121	0,121	0,126	0,124	0,123	0,124	0,125	0,121	0,124	0,123	0,124
Indicated	Total	Tonnes	41 179	35 442	76 621	124 008	5 188	29 617	147 488	306 301	70 687	209 214	134 163	279 864	693 929	95 980	174 081	227 595	367 337	864 993	1 941 844
(%) LREO	0,092	0,116	0,103	0,084	0,096	0,065	0,093	0,086	0,100	0,084	0,073	0,094	0,087	0,066	0,079	0,072	0,072	0,073	0,081
(%) HREO	0,077	0,046	0,063	0,054	0,029	0,081	0,056	0,057	0,037	0,050	0,073	0,063	0,058	0,078	0,062	0,074	0,089	0,078	0,067
(%) TREO	0,169	0,162	0,166	0,137	0,125	0,147	0,148	0,143	0,137	0,134	0,146	0,157	0,146	0,144	0,141	0,146	0,161	0,151	0,149
Grand Total	Tonnes	163 841	436 885	600 726	239 850	168 783	446 692	600 575	1 455 900	307 989	287 669	530 974	751 795	1 878 426	631 693	634 007	671 848	657 000	2 594 548	6 529 600
(%) LREO	0,069	0,067	0,067	0,074	0,061	0,063	0,068	0,067	0,073	0,078	0,065	0,084	0,076	0,072	0,076	0,068	0,066	0,070	0,071
(%) HREO	0,147	0,157	0,154	0,095	0,136	0,134	0,170	0,143	0,099	0,064	0,106	0,118	0,103	0,141	0,139	0,134	0,123	0,134	0,129
(%) TREO	0,216	0,223	0,221	0,168	0,197	0,197	0,239	0,209	0,172	0,141	0,171	0,202	0,179	0,213	0,214	0,202	0,189	0,204	0,200

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Table 104 LOM Annual Ore Mined by Measured and Indicated Resource Class

Resource Class	Material Type	Unit	PS	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Total
Measured	High Grade	Tonnes	313 598	611 489	613 343	986 796	429 561	182 853	31 335	3 315	0	0	0	0	0	0	3 172 288
		(%) LREO	0,067	0,066	0,078	0,081	0,077	0,081	0,092	0,083	0,000	0,000	0,000	0,000	0,000	0,000	0,076
		(%) HREO	0,231	0,249	0,189	0,227	0,206	0,173	0,093	0,088	0,000	0,000	0,000	0,000	0,000	0,000	0,217
		(%) TREO	0,298	0,315	0,266	0,308	0,283	0,254	0,185	0,171	0,000	0,000	0,000	0,000	0,000	0,000	0,293
	Low Grade	Tonnes	210 507	538 110	571 154	742 759	347 646	443 948	124 889	22 959	17 416	3	0	0	0	0	3 019 390
		(%) LREO	0,055	0,056	0,060	0,054	0,066	0,068	0,070	0,057	0,056	0,053	0,000	0,000	0,000	0,000	0,059
		(%) HREO	0,073	0,071	0,066	0,075	0,059	0,056	0,056	0,063	0,060	0,057	0,000	0,000	0,000	0,000	0,067
		(%) TREO	0,128	0,126	0,126	0,129	0,125	0,123	0,125	0,120	0,116	0,111	0,000	0,000	0,000	0,000	0,126
Measured	Total	Tonnes	524 105	1 149 599	1 184 497	1 729 555	777 206	626 801	156 223	26 274	17 416	3	0	0	0	0	6 191 678
		(%) LREO	0,062	0,061	0,069	0,069	0,072	0,072	0,074	0,061	0,056	0,053	0,000	0,000	0,000	0,000	0,068
		(%) HREO	0,168	0,166	0,129	0,162	0,140	0,090	0,063	0,066	0,060	0,057	0,000	0,000	0,000	0,000	0,144
		(%) TREO	0,230	0,227	0,198	0,231	0,212	0,161	0,137	0,127	0,116	0,111	0,000	0,000	0,000	0,000	0,211
Indicated	High Grade	Tonnes	27 962	69 852	183 964	216 597	670 439	698 947	1 244 962	588 588	1 081 672	1 007 992	827 259	541 195	1 139 446	406 976	8 705 853
		(%) LREO	0,140	0,117	0,120	0,077	0,086	0,090	0,084	0,107	0,092	0,098	0,096	0,114	0,088	0,111	0,094
		(%) HREO	0,091	0,088	0,088	0,157	0,196	0,170	0,247	0,100	0,139	0,194	0,152	0,090	0,116	0,093	0,157
		(%) TREO	0,231	0,205	0,208	0,234	0,282	0,259	0,331	0,207	0,231	0,292	0,248	0,204	0,204	0,204	0,251
	Low Grade	Tonnes	48 659	236 449	509 964	648 396	1 180 354	1 150 116	1 563 887	1 203 669	1 694 617	1 855 232	2 409 791	1 400 330	2 592 622	615 014	17 109 101
		(%) LREO	0,081	0,077	0,076	0,071	0,072	0,069	0,065	0,069	0,061	0,069	0,068	0,066	0,062	0,073	0,067
		(%) HREO	0,046	0,048	0,048	0,052	0,053	0,054	0,060	0,056	0,066	0,055	0,055	0,058	0,064	0,057	0,058
		(%) TREO	0,128	0,125	0,123	0,123	0,125	0,123	0,125	0,125	0,126	0,124	0,124	0,124	0,126	0,130	0,125
Indicated	Total	Tonnes	76 621	306 301	693 929	864 993	1 850 794	1 849 063	2 808 849	1 792 257	2 776 289	2 863 223	3 237 050	1 941 526	3 732 069	1 021 990	25 814 954
		(%) LREO	0,103	0,086	0,087	0,073	0,077	0,077	0,073	0,082	0,073	0,079	0,075	0,079	0,070	0,088	0,076
		(%) HREO	0,063	0,057	0,058	0,078	0,105	0,098	0,143	0,070	0,094	0,104	0,080	0,067	0,080	0,071	0,091
		(%) TREO	0,166	0,143	0,146	0,151	0,182	0,175	0,217	0,152	0,167	0,183	0,155	0,146	0,150	0,159	0,168
Grand Total	Tonnes	600 726	1 455 900	1 878 426	2 594 548	2 628 000	2 475 864	2 965 072	1 818 531	2 793 705	2 863 226	3 237 050	1 941 526	3 732 069	1 021 990	32 006 632
		(%) LREO	0,067	0,076	0,070	0,075	0,076	0,073	0,081	0,073	0,079	0,075	0,079	0,070	0,088	0,075	0,071
		(%) HREO	0,143	0,103	0,134	0,115	0,096	0,139	0,070	0,094	0,104	0,080	0,067	0,080	0,071	0,101	0,129
		(%) TREO	0,209	0,179	0,204	0,191	0,171	0,212	0,152	0,167	0,183	0,155	0,146	0,150	0,159	0,176	0,200

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First 3-Years Quarterly ROM Mined by Resource Class
Figure 151 ROM Mined by Resource Class

Annual ROM Mined by Resource Class

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Table 105 LOM Ex-Pit ROM to Destinations

Material Type	DESTINATION	Unit	PS	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Total
High-Grade	MILL /FLOATION	Tonnes		301 187	502 279	1 029 692	1 100 000	650 000	1 098 736	591 903	1 035 000	932 107	813 472	541 195	890 178	406 976	9 892 727
(%) LREO		0,065	0,073	0,077	0,082	0,074	0,078	0,107	0,090	0,096	0,095	0,114	0,075	0,111	0,086
(%) HREO		0,338	0,207	0,235	0,200	0,207	0,272	0,100	0,143	0,206	0,154	0,090	0,130	0,093	0,185
(%) TREO		0,403	0,279	0,313	0,282	0,281	0,350	0,207	0,233	0,302	0,249	0,204	0,206	0,204	0,271
High-Grade S/Pile	Tonnes	341 560	380 153	295 028	173 701		231 800	177 561		46 672	75 885	13 787		249 268		1 985 414
(%) LREO	0,073	0,077	0,113	0,100		0,126	0,122		0,139	0,128	0,133		0,131		0,104
(%) HREO	0,219	0,149	0,095	0,089		0,067	0,066		0,043	0,046	0,037		0,066		0,113
(%) TREO	0,292	0,226	0,208	0,189		0,193	0,189		0,181	0,174	0,171		0,197		0,217
High-Grade Total		Tonnes	341 560	681 340	797 308	1 203 392	1 100 000	881 800	1 276 297	591 903	1 081 672	1 007 992	827 259	541 195	1 139 446	406 976	11 878 141
(%) LREO	0,073	0,072	0,088	0,080	0,082	0,088	0,084	0,107	0,092	0,098	0,096	0,114	0,088	0,111	0,089
(%) HREO	0,219	0,232	0,165	0,214	0,200	0,170	0,244	0,100	0,139	0,194	0,152	0,090	0,116	0,093	0,173
(%) TREO	0,292	0,304	0,253	0,295	0,282	0,258	0,328	0,207	0,231	0,292	0,248	0,204	0,204	0,204	0,262
Low-Grade	MILL /FLOATION	Tonnes								134 250	149 250	991	909	302 725		177 465	765 589
(%) LREO								0,042	0,031	0,019	0,023	0,056		0,068	0,051
(%) HREO								0,084	0,097	0,109	0,112	0,071		0,067	0,077
(%) TREO								0,126	0,128	0,129	0,134	0,127		0,134	0,129
XRT SORTER	Tonnes		512 901	991 419	1 386 814	1 528 000	1 528 000	1 525 986	992 946	1 191 000	1 747 101	2 301 045	1 034 857	2 367 288	437 548	17 544 906
(%) LREO		0,053	0,064	0,062	0,071	0,068	0,063	0,070	0,058	0,068	0,068	0,066	0,060	0,075	0,065
(%) HREO		0,074	0,060	0,065	0,055	0,056	0,063	0,055	0,070	0,057	0,056	0,056	0,067	0,054	0,060
(%) TREO		0,127	0,125	0,126	0,125	0,124	0,126	0,125	0,127	0,125	0,124	0,122	0,127	0,128	0,125
Low-Grade S/Pile	Tonnes	259 167	261 658	89 700	4 341		66 064	162 790	99 432	371 783	107 143	107 837	62 748	225 334		1 817 996
(%) LREO	0,060	0,080	0,096	0,083		0,091	0,091	0,097	0,082	0,087	0,082	0,104	0,078		0,082
(%) HREO	0,068	0,044	0,025	0,021		0,027	0,031	0,026	0,041	0,032	0,032	0,024	0,041		0,040
(%) TREO	0,128	0,123	0,121	0,105		0,118	0,122	0,122	0,122	0,119	0,114	0,128	0,119		0,122
Low-Grade Total		Tonnes	259 167	774 559	1 081 119	1 391 155	1 528 000	1 594 064	1 688 775	1 226 628	1 712 033	1 855 234	2 409 791	1 400 330	2 592 622	615 014	20 128 491
(%) LREO	0,060	0,062	0,067	0,062	0,071	0,069	0,065	0,069	0,061	0,069	0,068	0,066	0,062	0,073	0,066
(%) HREO	0,068	0,064	0,057	0,065	0,055	0,055	0,060	0,056	0,066	0,055	0,055	0,058	0,064	0,057	0,059
(%) TREO	0,128	0,126	0,124	0,126	0,125	0,123	0,125	0,125	0,126	0,124	0,124	0,124	0,126	0,130	0,125
Total Mined		Tonnes	600 726	1 455 900	1 878 426	2 594 548	2 628 000	2 475 864	2 965 072	1 818 531	2 793 705	2 863 226	3 237 050	1 941 526	3 732 069	1 021 990	32 006 632
(%) LREO	0,067	0,067	0,076	0,070	0,075	0,076	0,073	0,081	0,073	0,079	0,075	0,079	0,070	0,088	0,075
(%) HREO	0,154	0,143	0,103	0,134	0,115	0,096	0,139	0,070	0,094	0,104	0,080	0,067	0,080	0,071	0,101
(%) TREO	0,221	0,209	0,179	0,204	0,191	0,171	0,212	0,152	0,167	0,183	0,155	0,146	0,150	0,159	0,176

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Table 106 LOM Stockpile Reclaim to Mill/Flotation

Material Type	DESTINATION	Unit	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Total
High Grade – Ex-HG S/Pile	FLOT_MILL	Tonnes	221 713	421 921	46 808		450 000	1 264	458 097		1 034	86 528	48 782		249 268	1 985 414
		(%) LREO	0,072	0,070	0,082		0,110	0,125	0,124		0,171	0,134	0,129		0,131	0,104
		(%) HREO	0,234	0,169	0,140		0,093	0,053	0,067		0,038	0,046	0,041		0,066	0,113
		(%) TREO	0,307	0,239	0,222		0,203	0,178	0,190		0,209	0,179	0,170		0,197	0,217
Low Grade – Ex XRT	FLOT_MILL	Tonnes	181 600	321 000	373 900	382 000	382 000	382 000	297 750	297 750	547 868	581 091	297 275	591 822	204 535	4 840 591
		(%) LREO	0,112	0,141	0,133	0,147	0,141	0,130	0,153	0,120	0,148	0,141	0,143	0,125	0,164	0,138
		(%) HREO	0,154	0,118	0,129	0,113	0,116	0,131	0,106	0,145	0,110	0,117	0,110	0,139	0,094	0,122
		(%) TREO	0,266	0,259	0,262	0,260	0,257	0,261	0,259	0,265	0,258	0,258	0,254	0,264	0,258	0,260
RECLAIM Total		Tonnes	403 313	742 921	420 708	382 000	832 000	383 264	755 847	297 750	548 902	667 619	346 056	591 822	453 803	6 826 006
		(%) LREO	0,090	0,100	0,128	0,147	0,124	0,130	0,135	0,120	0,148	0,140	0,141	0,125	0,146	0,128
		(%) HREO	0,198	0,147	0,130	0,113	0,104	0,131	0,082	0,145	0,110	0,108	0,101	0,139	0,079	0,119
		(%) TREO	0,288	0,247	0,258	0,260	0,228	0,261	0,217	0,265	0,258	0,248	0,242	0,264	0,225	0,248

Table 107 LOM Marginal Stockpiled

DESTINATIONS	LIFT	Unit	PS	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Total
MAR2_STK+	Lift-986	Tonnes						22 335	67 131	110 373	169 155	54 026	155 003	70			578 094
		(%) LREO						0,033	0,039	0,035	0,034	0,030	0,034	0,045			0,034
		(%) HREO						0,059	0,051	0,055	0,055	0,060	0,056	0,045			0,055
		(%) TREO						0,092	0,089	0,090	0,089	0,090	0,089	0,090			0,090
MAR4_STK+	Lift-1000	Tonnes	129 017	398 809	605 651	727 055	1 062 394	947 673	636 575	607 058	694 881	1 059 958	1 378 018	865 692	1 084 828	129 260	10 326 868
		(%) LREO	0,044	0,047	0,055	0,052	0,053	0,051	0,049	0,052	0,049	0,050	0,051	0,050	0,047	0,053	0,050
		(%) HREO	0,045	0,043	0,035	0,038	0,036	0,039	0,041	0,037	0,041	0,040	0,039	0,040	0,043	0,037	0,039
		(%) TREO	0,089	0,090	0,090	0,090	0,089	0,090	0,090	0,090	0,090	0,090	0,090	0,090	0,090	0,091	0,090
Marginal Total	Tonnes	129 017	398 809	605 651	727 055	1 062 394	970 008	703 706	717 431	864 036	1 113 984	1 533 021	865 762	1 084 828	129 260	10 904 963
		(%) LREO	0,044	0,047	0,055	0,052	0,053	0,051	0,048	0,050	0,046	0,049	0,049	0,050	0,047	0,053	0,050
		(%) HREO	0,045	0,043	0,035	0,038	0,036	0,039	0,042	0,040	0,044	0,041	0,041	0,040	0,043	0,037	0,040
		(%) TREO	0,089	0,090	0,090	0,090	0,089	0,090	0,090	0,090	0,090	0,090	0,090	0,090	0,090	0,091	0,090

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Table 108 LOM XRT Reject to A4N Waste Rock Dump

DESTINATION	LIFTS	Data	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Total
A4N WST	Lift-970	Tonnes	399 520	706 200												1 105 720
(%) LREO	0,027	0,034												0,032
(%) HREO	0,038	0,029												0,032
(%) TREO	0,065	0,063												0,064
Lift-990	Tonnes			822 580	840 400										1 662 980
(%) LREO			0,033	0,036										0,034
(%) HREO			0,031	0,028										0,030
(%) TREO			0,064	0,064										0,064
Lift-1010	Tonnes					840 400	840 400	655 050	473 370						2 809 220
(%) LREO					0,035	0,032	0,037	0,028						0,033
(%) HREO					0,028	0,032	0,026	0,037						0,030
(%) TREO					0,063	0,064	0,063	0,065						0,064
Lift-1030	Tonnes								181 680	1 205 310	1 230 262				2 617 251
(%) LREO								0,033	0,036	0,034				0,035
(%) HREO								0,032	0,027	0,029				0,028
(%) TREO								0,065	0,063	0,063				0,063
Lift-1050	Tonnes										48 140	654 005	1 302 008	449 977	2 454 129
(%) LREO										0,036	0,035	0,031	0,040	0,034
(%) HREO										0,030	0,027	0,034	0,023	0,030
(%) TREO										0,065	0,062	0,065	0,063	0,064
A4N WST Total	Tonnes	399 520	706 200	822 580	840 400	840 400	840 400	655 050	655 050	1 205 310	1 278 401	654 005	1 302 008	449 977	10 649 301
(%) LREO	0,027	0,034	0,033	0,036	0,035	0,032	0,037	0,029	0,036	0,035	0,035	0,031	0,040	0,034
(%) HREO	0,038	0,029	0,031	0,028	0,028	0,032	0,026	0,036	0,027	0,029	0,027	0,034	0,023	0,030
(%) TREO	0,065	0,063	0,064	0,064	0,063	0,064	0,063	0,065	0,063	0,063	0,062	0,065	0,063	0,064

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Table 109 LOM Waste to Waste Rock Dumps

Material Type	DESTINATION	LIFTS	PS	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Total
Min Waste	A2B_WST	Lift-970						41 886	85 541	131 023	101 635						360 085
Lift-990									116 522	69 083	198 215	88			383 908
A2B_WST Total						41 886	85 541	131 023	218 157	69 083	198 215	88			743 993
Waste	A4N_WST	Lift-970	2 513 192	8 640 003	13 370 158	1 436 677											25 960 029
Lift-990				13 358 351	10 772 958	11 736 721									35 868 030
		Lift-1010						913 134	8 061 055	10 559 553	7 520 714						27 054 456
Lift-1030										9 776 448	9 321 384				19 097 832
Lift-1050												4 652 499	4 184 210	184 225	9 020 935
A4N_WST Total	2 513 192	8 640 003	13 370 158	14 795 027	10 772 958	12 649 855	8 061 055	10 559 553	7 520 714	9 776 448	9 321 384	4 652 499	4 184 210	184 225	117 001 282
A4S_WST	Lift-990	1 557 065	2 205 289	3 045 764	4 028 304	9 259 648	2 435 279									22 531 349
Lift-1010						4 448 503	7 887 514	6 251 398	7 017 194	2 302 071					27 906 680
Lift-1030										6 801 065	4 062 932	8 039 442	2 334 408	27 827	21 265 675
A4S_WST Total	1 557 065	2 205 289	3 045 764	4 028 304	9 259 648	6 883 782	7 887 514	6 251 398	7 017 194	9 103 136	4 062 932	8 039 442	2 334 408	27 827	71 703 704
A2B_WST	Lift-970						501 604	3 820 112	3 845 064	2 960 732						11 127 512
Lift-990									2 120 462	897 123	1 647 397	683			4 665 666
A2B_WST Total						501 604	3 820 112	3 845 064	5 081 194	897 123	1 647 397	683			15 793 177
Waste Total	4 070 257	10 845 291	16 415 922	18 823 331	20 032 606	20 035 242	19 768 681	20 656 015	19 619 102	19 776 707	15 031 714	12 692 624	6 518 619	212 052	204 498 163

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Table 110 LOM Annual Mill/Flotation Feed

MATERIAL FROM	Unit	Y01	Y02	Y03	Y04	Y05	Y06	Y07	Y08	Y09	Y10	Y11	Y12	Y13	Total
BOS PCP	Tonnes	181 600	321 000	373 900	382 000	382 000	382 000	297 750	297 750	547 868	581 091	297 275	591 822	204 535	4 840 591
	(%) LREO	0,112	0,141	0,133	0,147	0,141	0,130	0,153	0,120	0,148	0,141	0,143	0,125	0,164	0,138
	(%) HREO	0,154	0,118	0,129	0,113	0,116	0,131	0,106	0,145	0,110	0,117	0,110	0,139	0,094	0,122
	(%) TREO	0,266	0,259	0,262	0,260	0,257	0,261	0,259	0,265	0,258	0,258	0,254	0,264	0,258	0,260
EX PIT	Tonnes	301 187	502 279	1 029 692	1 100 000	650 000	1 098 736	726 153	1 184 250	933 098	814 381	843 920	890 178	584 442	10 658 316
	(%) LREO	0,065	0,073	0,077	0,082	0,074	0,078	0,095	0,082	0,096	0,095	0,093	0,075	0,098	0,084
	(%) HREO	0,338	0,207	0,235	0,200	0,207	0,272	0,097	0,137	0,206	0,154	0,083	0,130	0,085	0,177
	(%) TREO	0,403	0,279	0,313	0,282	0,281	0,350	0,192	0,220	0,302	0,249	0,176	0,206	0,182	0,261
HG STK	Tonnes	221 713	421 921	46 808		450 000	1 264	458 097		1 034	86 528	48 762		249 268	1 985 414
	(%) LREO	0,072	0,070	0,082		0,110	0,125	0,124		0,171	0,134	0,129		0,131	0,104
	(%) HREO	0,234	0,169	0,140		0,093	0,053	0,067		0,038	0,046	0,041		0,066	0,113
	(%) TREO	0,307	0,239	0,222		0,203	0,178	0,190		0,209	0,179	0,170		0,197	0,217
Grand Total	Tonnes	704 500	1 245 200	1 450 400	1 482 000	1 482 000	1 482 000	1 482 000	1 482 000	1 482 000	1 482 000	1 189 977	1 482 000	1 038 245	17 484 322
	(%) LREO	0,079	0,089	0,092	0,099	0,102	0,091	0,115	0,090	0,115	0,115	0,107	0,095	0,119	0,101
	(%) HREO	0,258	0,171	0,205	0,178	0,149	0,236	0,090	0,139	0,170	0,133	0,088	0,134	0,082	0,155
	(%) TREO	0,337	0,260	0,297	0,276	0,251	0,327	0,205	0,229	0,285	0,248	0,195	0,229	0,201	0,256
	(ppm) La2O3	197,566	227,329	239,580	259,888	266,776	228,641	321,257	234,058	302,165	304,912	281,211	236,231	319,529	264,708
	(ppm) Ce2O3	356,088	404,896	421,453	454,589	471,806	413,972	531,313	405,591	529,000	523,086	497,206	436,802	554,731	464,151
	(ppm) Pr2O3	38,497	43,501	44,009	47,542	49,794	44,249	54,200	42,526	55,181	54,990	52,267	47,830	57,447	48,861
	(ppm) Nd2O3	143,713	161,299	159,177	171,520	179,984	165,186	192,028	157,130	199,736	202,293	186,719	176,579	201,521	177,592
	(ppm) Sm2O3	57,314	55,257	53,663	54,234	56,250	62,154	54,766	56,516	63,536	68,711	53,286	54,646	55,043	57,498
	(ppm) Eu2O3	28,321	24,146	24,283	23,490	22,901	28,848	20,309	23,956	25,769	25,961	20,521	20,833	18,041	23,623
	(ppm) Gd2O3	131,752	104,879	110,816	102,608	96,876	129,250	77,242	98,984	108,318	103,373	78,664	86,990	68,723	99,523
	(ppm) Tb2O3	32,629	24,019	26,817	23,571	21,031	30,444	14,786	20,935	23,460	20,499	14,532	18,905	12,941	21,725
	(ppm) Dy2O3	233,568	162,623	188,422	162,018	139,344	212,505	91,120	136,656	158,279	130,986	89,245	128,493	81,394	145,803
	(ppm) Ho2O3	52,303	35,043	41,362	35,111	29,599	46,103	17,928	27,967	33,390	26,317	17,570	27,335	16,367	30,863
	(ppm) Er2O3	157,454	101,192	122,191	102,633	85,971	138,012	50,290	80,199	97,957	74,873	49,334	80,515	47,459	90,082
	(ppm) Tm2O3	24,120	15,027	18,436	15,415	12,870	20,859	7,129	11,647	14,482	10,824	6,858	11,847	6,837	13,350
	(ppm) Yb2O3	146,579	90,059	112,112	92,254	77,804	126,674	43,600	70,215	87,211	65,471	41,759	74,049	42,090	80,978
	(ppm) Lu2O3	21,626	13,007	16,289	13,348	11,235	18,253	6,207	9,957	12,554	9,413	5,899	10,824	6,164	11,697
	(ppm) Y2O3	1 749,538	1 140,863	1 387,149	1 206,445	992,484	1 604,982	566,604	908,721	1 141,987	860,766	558,560	877,022	521,225	1 027,352

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16.12.5. Haul Network and Cycle Times

The MP Haulage application was used to set-up the haulage network to be used in MPSO. The cycle times were determined in MPSO from all 3 003 mining cuts to all possible destinations. Furthermore, the cycle times for all stockpiles' reclaiming to various possible destinations and in-lift haul distances and cycle times are calculated to simulate the dump level- and height progression during the schedule.

The haulage profiles in-pit and on surface is illustrated in Figure 152 below. The total cycle time detail is too voluminous to show in the report and is available as a separate text file if required.

Figure 152 Schedule Haulage Setup

16.12.6. Yearly Progress Plots

The annual progress plots for Pit A2B and Pit A4 are illustrated in Figure 153 through Figure 172.

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Figure 153 Year-00 (Pre-Strip) Pit A4 Progress

Figure 154 Year-01 Pit A4 Progress

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Figure 155 Year-02 Pit A4 Progress

Figure 156 Year-03 Pit A4 Progress

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Figure 157 Year-04 Pit A4 Progress

Figure 158 Year-05 Pit A4 Progress

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Figure 159 Year-05 Pit A2B Progress

Figure 160 Year-06 Pit A4 Progress

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Figure 161 Year-06 Pit A2B Progress

Figure 162 Year-07 Pit A4 Progress

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Figure 163 Year-07 Pit A2B Progress

Figure 164 Year-08 Pit A4 Progress

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Figure 165 Year-08 Pit A2B Progress

Figure 166 Year-09 Pit A4 Progress

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Figure 167 Year-09 Pit A2B Progress

Figure 168 Year-10 Pit A4 Progress

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Figure 169 Year-10 Pit A2B Progress

Figure 170 Year-11 Pit A4 Progress

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Figure 171 Year-12 Pit A4 Progress

Figure 172 Year-13 Pit A4 Progress

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17. Recovery Methods

The proposed flow sheet for the Lofdal process plant is based on testwork results as reported in Chapter 13. The throughput capacity for this plant is 3.01 million tonne per annum of run of mine ore for crushing and ore sorting and 1.482 million tonne per annum for further processing.

The plant will consist of three main sections: the crushers and sorters, the concentrator and the refinery.

The crushers and sorters will comprise ROM feed reception, low grade ore primary and secondary crushing, ore sorting and tertiary crushing and high grade ore primary, secondary and tertiary crushing.

The concentrator will comprise crushed ore stockpile, milling, and flotation. The flotation circuit will include roughers, cleaners, concentrate thickening and filtration, tails thickening and transfer of the tailings underflow to the tailings storage facility. The water recovered from the tailing and concentrate overflow thickeners will be pumped into the front-end process water storage facility. The concentrate cake will be transferred to the refinery section.

The refinery section of the plant will be dedicated to the extraction, purification, and precipitation of REEs through a series of operations such as sulphation roast, water leach, impurities removal and REE precipitation. The REE will be precipitated as a carbonate, dried and packaged for sale.

The ROM contains an appreciable amount of uranium which will be removed from solution by ion exchange and precipitated as ammonium di-uranate (ADU) using ammonium hydroxide. This precipitate will be thickened and the ADU slurry will be sent to a Namibian uranium producer.

Water will be recycled via the back-end process water tank where possible and effluent streams from this section will be pumped into the tailings transfer tank.

Reagent handling, utilities and services will be installed and will consist of the following:

Flotation reagents
Flocculants and coagulants
Sodium hydroxide
Sodium silicate
Sulphuric acid
Magnesium carbonate
Sodium carbonate
Ammonium hydroxide
Hydrogen peroxide
Plant Air
Process water
Raw water
Potable water
Gland seal water
Fire Water
Cooling water
Demineralized water
Boiler

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The process is depicted in the block flow Diagram (Figure 173).

Figure 173 Block Flow Diagram Processing Plant

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17.1. Process Design Basis

The specific area codes as defined by the project work breakdown structure (WBS) are given in Table 111. The process block flow diagram is shown in Figure 173 with all functional areas included.

Table 111 Area Codes According to Project WBS

Code	Description
B02	Crushing and sorting
C01	Primary Mill and Stockpile
C02	Classification
D01	Rougher flotation
D02	Cleaner flotation
E01	Tailings thickening and pumping
E02	Neutralization
F01	Concentrate thickening
G01	Sulphation roast
G02	Leach
G03	Impurity precipitation and filtration
H01	Uranium IX
H02	ADU precipitation
H03	ADU Dispatch
J01	RE carbonate precipitation and thickening
M02	REE carbonate drying and packaging
N01	Hydro reagents
N02	Reagents
P01	Utilities
P02	Water systems

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17.2. Process Design Criteria, Summary

Process design criteria (PDC) is compiled to provide guidance in terms of critical process design input, outputs and requirements associated with the global control and operational philosophy. Table 112 below presents the major design parameters used in the development of the Mass Balance for process design of a range of operating points. The source of parameters and conditions adopted in the PDC includes:

Client – Namibia Critical Metals
Testwork results performed by Mintek, SGS Lakefield, Geolabs SA, Rados International, Light Deep Earth Laboratories (LDE), IMS Engineering (Steinert), Tomra.
Assumptions (by SGS Bateman) based on previous similar projects
Recommendations from suppliers (vendors)

Table 112 Major Design Parameters
AREA – General

Description	Units	Value
Feed Tonnage	tpa	3 010 000
Solids Density	SG	2.66
Feed % Solids	%	98
High Grade Feed Grade	% TREO	0.262
Low Grade Feed Grade	% TREO	0.125

AREA – High Grade Crushing

Description	Units	Value
Primary Crushing
Crushing Availability	%	68
Crusher	Type	Jaw
Crusher Feed Method	Type	Front End Loader
Operation	Type	Open Circuit
Product Size P80	mm	79
Secondary Crushing
Crusher	Type	Cone
Feed method	Type	Vibrating screen
Operation	Type	Closed Circuit
Product Size P80	mm	24
Tertiary Crushing
Crusher	Type	VSI
Feed method	Type	Vibrating screen
Operation	Type	Closed Circuit
Product Size P100	mm	17

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AREA – Low Grade crushing and sorting

Description	Units	Value	Remarks
Primary Crushing
Crushing Availability	%	68
Crusher	Type	Jaw
Crusher Feed Method	Type	Front End Loader
Operation	Type	Open Circuit
Product Size P80	mm	90
Secondary Crushing
Crusher	Type	Cone
Feed method	Type	Vibrating screen
Operation	Type	Closed Circuit
Product Size P80	mm	58
Ore sorting
Sorter	Type	XRT
Feed method	Type	FEL and Conveyor
Coarse Ore Sorting	mm	50 – 25
Fine Ore Sorting	mm	25 - 10
Fines discard	mm	10-0
Tertiary Crushing
---	---	---	---
Crusher	Type	VSI
Feed method	Type	Vibrating screen
Operation	Type	Closed Circuit
Product Size P80	mm	17

AREA– Grinding

Description	Units	Value
Ball milling
Duty		1 stage Ball Milling
Operating mode	Type	closed circuit with scalping and dewatering cyclones
Grinding Mode	Type	Wet
Availability	%	90
Mill circuit feed particle size, F100	mm	20.00
Mill circuit feed particle size, F80	mm	12.00
Closed circuit recirculation load (@ mill feed)	%	250.00
Milling circuit product, P80	μm	38
Bond ball mill Work-Index	kWh/t	16.70
Closed Circuit Classification
Classifier Type	Type	Flat Bottom Cyclones
Final Cyclone Overflow P80	μm	43

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AREA– Flotation

Description	Units	Value	Remarks
Flotation
Flotation Conditioning	Type	High Intensity	1 800 rpm
Flotation Conditioning % Solids	%	35
Flotation Feed Density	SG	1.3
Overall Float Mass Pull	%	3.78
Overall Float TREO Recovery	%	52-68
Primary Collector addition rate	g/t Feed	1 235	Florrea 3900Z
Secondary Collector addition rate	g/t Feed	260	Florrea 3000
Depressant addition rate	g/t Feed	213	Sodium Silicate
Dispersant addition rate	g/t Feed	46	Calgon
Float Concentrate Thickening
Thickener Underflow Density	% solids	50
Thickener type	Type	Conventional
Flocculant addition - current on normal ore	g/t Feed	42
Rise Rate (design)	m/day	1 861

AREA– Water Leach and Acid Bake

Description	Units	Value	Remarks
Water Leach
H_{2}SO_{4} Dosage	kg/t conc.	1 500
Acid Bake temperature	°C	300 to 600	Allowance for possible 2 stage Sulphation roasting.
Acid Bake residence time	h	1.5
Water Leach Solid Density	% solids	16 -18.4
Number of Tanks	#	4	Including quench tank
Total Residence Time	hours	1.00
Water leach temperature	°C	50

AREA– Impurity Removal – Primary Precipitation

Description	Units	Value
-
Precipitant		MgCO_{3}
MgCO_{3} Dosage	kg/t conc.	707
MgCO_{3} Reagent Strength	% (w/w)	20

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AREA– Secondary Precipitation

Description	Units	Value

Precipitant		MgCO₃
MgCO₃ Dosage	kg/t conc.	42.3
MgCO₃ Reagent Strength	%	20
Number of Tanks	#	4
Total Residence Time (Required)	minutes	60

AREA– U IX and Precipitation

Description	Units	Value
Uranium Ion Exchange
U IX % Recovery	%	99.9
U Tenor in Eluate	g/l	5
Eluate Acid Concentration	g/l	100
Uranium Precipitation
Precipitant		NH₄OH
Precipitant Dosage		pH Controlled
pH		7.3

AREA– REE Precipitation

Description	Units	Value
REE Precipitation
Precipitant		Na₂CO₃
Precipitant Dosage	kg/t conc.	34.36
Number of Tanks	#	1.0
Total Residence Time (Required)	minutes	60.0
Re-seed %	%	350
Target pH	pH	6.75

Overall REE Recoveries

	Overall Process Plant Recovery [%]
Total REO	59.4
Heavy REO	62.2

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17.3. Plant Description

This process description should be read in conjunction with the following process flow diagrams (drawings):

Drawing Number	Title
B02013P1200010001 Rev G	Front End Crushing and Sorting
C01013P1200010001 Rev G	Crushed Ore Stockpiling & Primary Milling
C02013P1200010001 Rev G	Classifying Hydro-cyclone
D01023P1200010001 Rev F	Rougher Flotation
D02023P1200010001 Rev F	Cleaners Stage 1 Flotation
D02043P1200010001 Rev G	Cleaners Stage 2 Flotation
D02053P1200010001 Rev F	Flotation Air Blowers
E01013P1200010001 Rev F	Tailings Thickening
E01023P1200010001 Rev F	Tailings Pumping
E02013P1200010001 Rev F	Neutralization Back-End
F01013P1200010001 Rev G	Concentrate Thickening
F21023P1200010001 Rev F	Concentrate Filtration
G01013P1200010001 Rev G	Sulphation Roast
G02013P1200010001 Rev F	Leach
G03013P1200010001 Rev G	Impurities Precipitation and Filtration
G03013P1200010002 Rev F	Impurities Precipitation Thickening - Second Neutralization
H01013P1200010001 Rev F	Uranium Ion Exchange
H02013P1200010001 Rev F	ADU Precipitation and Thickening
H03013P1200010001 Rev F	ADU Slurry Product and Handling
J01013P1200010001 Rev G	REE Precipitation, Thickening and Filtration
M02013P1200010001 Rev F	REE Carbonate Drying and Packaging
N01023P1200010001 Rev G	Sulphuric Acid
N01053P1200010001 Rev G	Magnesium Carbonate
N01063P1200010001 Rev F	Sodium Carbonate
N01073P1200010001 Rev F	Ammonium Hydroxide
N01093P1200010001 Rev F	Lime
N02023P1200010001 Rev F	Primary & Secondary Collector
N02033P1200010001 Rev F	Flocculant
N02043P1200010001 Rev F	Sodium Hydroxide
N02053P1200010001 Rev F	Calgon
P01013P1200010001 Rev F	Utilities – Plant and Instrument Air
P02013P1200010001 Rev F	Process Water

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Drawing Number	Title
P02023P1200010001 Rev F	Raw Water
P02033P1200010001 Rev G	Potable Water
P02043P1200010001 Rev F	Gland Seal Water
P02053P1200010001 Rev F	Fire Water
P02063P1200010001 Rev F	Demineralized Water
P02073P1200010001 Rev G	Cooling Water Package

17.3.1. Concentrator

17.3.1.1. Ore Reception Crushing and Sorting

Run-of-mine ore is delivered by mine truck to high grade and low-grade stockpiles for the respective crushing circuits.

A front-end loader is used to withdraw high grade ore from the stockpile and feed the primary jaw crusher hopper. Prior to crushing, a vibrating grizzly screen removes undersize material. The jaw crusher operates in an open circuit.

The product from the primary crusher is conveyed to the triple deck secondary vibrating screen so that only the oversize material is fed to the secondary and tertiary crusher. The product from the secondary crusher recycled in closed circuit to the secondary vibrating screen. The product from the tertiary crusher recycled in closed circuit to the tertiary vibrating screen. The undersize is conveyed to the crushed ore stockpile.

A front-end loader is used to withdraw low grade ore from the stockpile and feed the primary jaw crusher hopper. Prior to crushing, a vibrating grizzly screen removes undersize material. The jaw crusher operates in an open circuit.

The product from the primary crusher is conveyed to the triple deck secondary vibrating screen so that only the oversize material is fed to the secondary crusher. The product from the secondary crusher recycled in closed circuit to the secondary vibrating screen. Sized product from the screen is conveyed to either a coarse (-50mm+25mm) or mids (-25+10mm) stockpile for or sorting or a fines (-10mm) discard stockpile.

The coarse and mids stockpiles are fed to the XRT ore sorters by front end loader. Sorter product is conveyed to the crushed ore stockpile. Sorter rejects are conveyed to a rejects stockpile.

The fine discard and reject stockpiles are loaded into the returning mine truck and discarded on the waste rock dumps by the mining contractor.

17.3.1.2. Milling (Ball Milling)

The ball mill circuit consists of a single primary mill. The mill is designed to produce the fine product sizing required for flotation, $P80 = 38$ microns. The primary ball mill operates in a closed circuit with a cyclone cluster. The cyclone overflow will feed the conditioning tank in the rougher flotation bank.

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17.3.1.3. Flotation

The flotation circuit will operate as a single train with rougher, cleaner and re-cleaner float banks. The re-cleaner concentrate will form the final product feeding the downstream hydrometallurgical sections. The rougher and cleaner tails will report to tailings. The flotation sections will be configured as follows:

Roughers: 5 x 70 m³ tank cells
Cleaners: 3 x 20 m³ tank cells
Re-cleaners: 2 x 5 m³ tank cells

A float tailings thickener will be used to de-water the final float tailings to the TSF, which is fed from the final tailings tank. A float concentrate thickener will be used to produce a thickener underflow product feeding forward to the hydrometallurgical processes.

17.3.1.4. Concentrate Dewatering and Filtration

The flotation concentrate from the cleaners is pumped into the concentrate thickener where it will be mixed with flocculant to enhance the solids settling rate. The concentrate thickener underflow contains 50% solids and is further filtered through a filter press. The filter cake produced contains 15% moisture. This cake is transferred to the REE refinery for further processing. The thickener overflow and the filtrate are pumped into the process water storage.

17.3.1.5. Tailings Dewatering and Transfer

The tailings thickener receives feed from the rougher flotation tailings tank. The thickener feed is diluted and mixed with flocculant to enhance the solids settling rate. The thickener produces an underflow stream containing 50% solids. The tailings thickener underflow is pumped to the tailings storage tank, where it is combined with a neutralisation slurry. The thickener overflow is collected and pumped to a process water dam. The tailings in the tailings storage tank are subsequently transferred to the tailings storage facility (TSF).

17.3.2. Refinery

17.3.2.1. Sulphation Roast

Flotation concentrate cake will report to acid mixing where concentrated sulphuric acid will be added to the cake in a pug mill prior to feeding the acid-bake kiln.

The design approach allowed for two stage sulphation roasting as the testwork was still underway at the time of design. The kiln will operate at 600°C. With a one-hour heating zone and one and a half hours at the target reaction temperature. The kiln will be specified in terms of MOC for sulphuric acid at the operating temperature. The kiln promotes the sulphation of the feed material. The final kiln product is cooled to a temperature which will allow the subsequent water leach process to operate at 50°C.

During the sulphation roast, REE elements react with sulphuric acid to form stable soluble rare earth sulphates while minimizing the formation of soluble impurity compounds (e.g., Fe, Al etc.). If the impurities are dissolved, they will be removed from the solution by precipitation after water leach.

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Y₂O₃	+ 3 H₂SO₄	= Y₂(SO₄)₃	+ 3 H₂O	(1)
La₂O₃	+ 3 H₂SO₄	= La₂(SO₄)₃	+ 3 H₂O	(2)
Pr₂O₃	+ 3 H₂SO₄	= Pr₂(SO₄)₃	+ 3 H₂O	(3)
Nd₂O₃	+ 3 H₂SO₄	= Nd₂(SO₄)₃	+ 3 H₂O	(4)
Sm₂O₃	+ 3 H₂SO₄	= Sm₂(SO₄)₃	+ 3 H₂O	(5)
Tb₂O₃	+ 3 H₂SO₄	= Tb₂(SO₄)₃	+ 3 H₂O	(6)
Dy₂O₃	+ 3 H₂SO₄	= Dy₂(SO₄)₃	+ 3 H₂O	(7)
Tm₂O₃	+ 3 H₂SO₄	= Tm₂(SO₄)₃	+ 3 H₂O	(8)
Fe₂O₃	+ 3 H₂SO₄	= Fe₂(SO₄)₃	+ 3 H₂O	(9)
Al₂O₃	+ 3 H₂SO₄	= Al₂(SO₄)₃	+ 3 H₂O	(10)

Sulfur rich off gas from the kiln is scrubbed using concentrated Sulphuric acid to recover the Sulphuric acid and recycle the acid back to the pug mil.

17.3.2.2. REE Water Leach

Solids mainly from the Acid Bake kiln are re-slurried with water in a four-tank leach operation. Secondary neutralization precipitates are recycled to the water leach for re-dissolution. This is done to selectively dissolve REEs from the kiln product. Leach efficiencies are expected to be high, supported by test work.

17.3.2.3. Impurities Precipitation and Filtration

The water leach slurry is then mixed with magnesium carbonate in the primary impurity removal stage followed by secondary neutralisation. The impurity removal stage selectively precipitates most of the dissolved thorium and a vast majority of the dissolved iron. Rare earth losses at this stage are expected to be low. The pulp from impurity removal will be thickened and filtered. Solids will report to final neutralization for discard to tailings while the thickener overflow and filter filtrate report forward in the circuit to second neutralization where further impurity removal will be carried out. The pH will be increased between 4 and 5 to precipitate iron, aluminium, thorium etc while minimizing the REEs losses. The resultant slurry will be thickened and recycled back to the water leach to recover any REE which might have coprecipitated with iron. The second neutralization thickener overflow transferred to uranium Ion Exchange (IX).

17.3.2.4. Uranium Ion Exchange

Uranium is selectively adsorbed from the impurity removal product solution using a fixed bed ion exchange circuit. The resin used will be a strong base type to facilitate uranium adsorption. Uranium recovery to the resin can be considered quantitative with negligible REE losses. Barren liquor proceeds down the circuit, containing the REE component. The resin is eluted with $100\mathrm{g / l}$ $\mathsf{H}_2\mathsf{SO}_4$ , producing an eluate of approximately $5\mathrm{g / l}$ $\mathsf{U}_3\mathsf{O}_8$ equivalent.

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17.3.2.5. ADU Precipitation

Eluate from ion exchange is precipitated using NH₄OH solution under pH control to produce an ADU precipitate product. This is thickened, washed and shipped as a final uranium product.

17.3.2.6. Rare Earth Elements Precipitation and Thickening

Uranium IX barren liquor is mixed with sodium carbonate under pH control to produce a rare earth carbonate precipitate.

The product from this stage is thickened, with thickener overflow reporting the final neutralization and tailings. Thickener underflow will be filtered.

Spillages within this area are directed into the REE precipitation tank.

17.3.2.7. Rare Earth Elements Carbonate Precipitate Drying and Packaging

REE cake will be conveyed from the filter to the dryer to reduce the free moisture of the REE precipitate cake. Residence time in the dryer will be controlled by varying the speed of the dryer drive. Dried REE precipitate cake will exit the dryer rotating drum to a bucket elevator then will be stored in the final product bin.

The dryer with all the accessories, which includes the burner, baghouse, fan, stack, packaging etc. will be a vendor package.

17.3.3. Reagents

17.3.3.1. Sulphuric Acid

Concentrated sulphuric acid (98%) will be delivered by road tankers, off-loaded into the carbon steel storage tanks by an offloading pump. The two tanks' overflow nozzles will be connected with an equalising line and the offloading pump will be interlocked with the high-level alarm in both tanks, providing for operator safety. There will be a vent on each tank connected to a common desiccant dryer to avoid moisture ingress into the tank.

Acid will be distributed from the storage tanks to the users, sulphation roast and SX, by the sulphuric acid supply pumps.

Sulphuric acid spillage from the offloading area is pumped to the neutralisation area by the spillage pump. A safety shower will be provided in this area in case of emergency.

17.3.3.2. Magnesium Carbonate

Magnesium carbonate (MgCO₃) will be purchased from suppliers and delivered to the site in bulk 1000 kg bags.

The bag will be hoisted and fed to the make-up tank through a hopper fitted with a bag breaker. The make-up tank, which will be fitted with an agitator, will be filled with a pre-determined volume of water to make a magnesium carbonate solution of 20% w/w of MgCO₃.

Magnesium carbonate will be added to the agitated make-up tank. The agitation will continue until the solids are dissolved. Further water will be added to adjust the solution to the required

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normality. This solution will be transferred to the dosing tank from where it will be dosed into the impurities removal (precipitation) tank.

The pH in the precipitation tank will be controlled by the addition of magnesium carbonate.

Spillage from the sodium carbonate make-up area will be pumped by the spillage pump the magnesium carbonate make-up tank.

17.3.3.3. Sodium Carbonate

Sodium carbonate (Na₂CO₃) will be purchased and delivered to the site in bulk 1000 kg bags.

The bag will be hoisted and fed to the make-up tank through the hopper fitted with a bag breaker. The make-up tank, which will be fitted with an agitator, will be filled with a predetermined volume of water to make a sodium carbonate solution of 20% w/w of Na₂CO₃.

Sodium carbonate will be added to the agitated make-up tank. The agitation will continue until the solids are dissolved. Further water will be added to adjust the solution to the required normality. This solution will be transferred to the dosing tank from where it will be pumped to the REE precipitation tank.

Sodium carbonate will also be used for acid bake kiln off gas scrubber solution make-up.

Spillage from the sodium carbonate make-up area will be pumped by the spillage pump the sodium carbonate make-up tank.

17.3.3.4. Ammonium Hydroxide

Ammonium hydroxide will be used to precipitate uranium as uranium diuranate.

Ammonium hydroxide will be delivered on site in containers and transferred to the storage tank where it will be mixed with water to make a 20% solution. The make-up system will consist of make-up tank, transfer pumps, dosing tank and dosing pumps. The make-up tank will be installed and fitted with an agitator.

17.3.3.5. Milk of Lime (Calcium Hydroxide)

Hydrated lime (Ca(OH)₂) will be purchased from local suppliers and transported by road to site in 1 m³ bulk bags.

The lime will be loaded into a hydrated lime mixing tank. The hydrated lime will be mixed with process water in the mixing tank to make up milk of lime slurry. The milk of lime will be transferred by transfer pumps to the milk of lime storage tank.

From the storage tank, the milk of lime will be pumped to neutralisation section via a ring main system returning to the tank.

Spillage from the area will be pumped by the spillage pump into the lime mixing tank.

17.3.3.6. Primary (3900z) and Secondary (3000) Collectors

The primary and secondary collectors are delivered to site stored in bags and are subsequently offloaded into their designated bag breakers. These collectors are then made up into solutions by mixing them with water within their respective mixing tanks. The collector solutions are then pumped to their individual collector dosing tanks from where each collector's dosing pumps

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distribute them to the flotation area for selective hydrophobic particles' attachment to the air bubbles.

17.3.3.7. Flocculant

Flocculant will be used in various sections of the plant as an aid to thickeners and filter performance.

The flocculant make-up systems will consist of bag breaker, hopper, feeder, mixing tanks, storage tanks, transfer pumps, positive displacement dosing pumps (one pump for each end user) and dilution mixers.

Flocculant powder will be supplied in 25 kg bags which will be manually loaded into the hopper. Raw water will be used to dissolve the flocculant in the mixing tank.

Flocculant spillage within the area will be collected in the sump and pumped to tailings neutralisation.

17.3.3.8. Sodium Hydroxide

Bulk caustic lye (NaOH) will be purchased and transported to the plant where it will be mixed with water to make a caustic solution.

This solution will be used for pH adjustment. The make-up tank will be installed with the agitator.

Spillage from the caustic make-up area will be pumped by the spillage pump into the caustic make-up tank.

17.3.3.9. Calgon

The Calgon is off-loaded and pumped to the Calgon distribution tank. The transfer between Calgon isotainer and the distribution tank is done using an off-loading pump. It is then pumped and distributed to the flotation area to strengthen the surface tension of air bubbles.

17.3.3.10. Sodium Silicate

Sodium silicate is delivered to site stored in bags. It is then made up into a solution by mixing it with water inside the mixing tank. The sodium silicate solution is then pumped to the dosing tank from where a dosing pump distributes it to the flotation area.

17.3.4. Utilities

17.3.4.1. Plant Air and Instrument Air (Compressed Air)

Compressed air for plant and instrument use will be provided by two compressors (one running and one standby). The compressor generates compressed air at 7 barg pressure and deliver to air receiver.

Plant air will be drawn directly from the first air receiver. The compressors will be automatically controlled by demand.

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High-quality instrument air will be required and drawn from the second air receiver and passed through a desiccant type of dryer (one running, one standby). There will be also filters before the dryers (for oil removal) and after the dryers (for fine particle removal).

17.3.4.2. Process Water

Process water will be used in various sections of the plant. Due to water scarcity, the discharge of water from the process into the environment will be minimised and treated where possible. The input of fresh raw water must be minimised. This will be achieved by using recycled water for general process duties such as wash-down, line flushing, mill feed dilution etc.

Process water will be stored in a dedicated process water pond, from where it will be pumped to the various users by raw water supply pumps. The concentrator and refinery sections have separate process water circuits.

17.3.4.3. Raw Water

The Lofdal plant will draw its raw water from a raw water pond via the raw water supply pumps.

17.3.4.4. Potable Water

A potable water treatment plant will be supplied with filtered raw water. The treatment plant is a vendor package. Potable water product will be stored in a potable water tank from which the supply pumps will deliver potable water to the various end-users, including the plant safety showers and top-up for the fire water storage tank.

To ensure a guaranteed uninterrupted supply of water to the safety showers, the potable water pumps are backed up with emergency power.

17.3.4.5. Gland Seal Water

Filtered raw water will be supplied to the gland seal water tank from which it will be pumped to the plant header feeding the various slurry pumps on the plant.

A separate high-pressure seal water system will be required to supply the tailings pumps. HP gland seal water supply pumps draw water from the main shared seal water tank.

17.3.4.6. Fire Water

The fire water system consists of a fire water storage tank supplying the fire water pumping. The pumping station will include a two main pumps. The first will be electrically driven and the second will be diesel driven. There will be also jockey pump to maintain system pressure and ensure that the main pumps do not start unnecessarily.

17.3.4.7. Demineralized Water

Demineralised water will be used on some processes (e.g., Uranium IX) to avoid introducing impurities which might compromise the quality of the product.

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17.3.4.8. Steam

The steam will be used to heat the process liquor to the operating temperature where it is required.

The steam condensate will be collected in the condensate tank and returned to the boiler for the make-up. The boiler system will be a vendor package.

17.3.4.9. Fuel

The diesel fuel storage facility will be completed by a third party. However, for the process plant, a provision has been made for a day tank within the area where the fuel is required. The diesel fuel is used in the following plant areas:

Fuel for mobile transport, particularly forklift trucks.
At the sulphation roast kiln, for start-up and supplementary heating.
At the REE carbonate precipitation dryer, for heating.
At the fire water pump station for the diesel-driven pump (1 m³ tank); and
The emergency power diesel generators.

The emergency power generators will start up automatically in the event of a plant power failure. They will supply power to all the critical plant items that would otherwise be severely impacted by a power outage – such as slurry tank agitators, certain slurry pumps, thickener rake drives and potable water pumps.

17.4. Safety and Risk Assessment

17.4.1. Background and Objectives

Project risks were identified during the study. These risks were rated and ranked in terms of likelihood and consequence including mitigation action plans. This risk register will be carried forward, monitored, and expanded during project execution.

17.4.2. Risk Management Process

The following risk management framework was followed for project team risk reviews:

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Communicate and Consult - Communication of risk registers and response plans will be communicated with all identified stakeholders.

Establish the Context - The context for risk identification will be used to divide the project risks into different categories. This will ensure that the correct audience participates in the risk identification and that the most appropriate resources are allocated to risk response plans.

Risk Identification - The following techniques will be used to identify project risks:
- Brainstorming
- Individual risk registers by discipline,
- Risk workshops
- Incident investigation (especially during construction)
- Criticality reviews.

Risk Analysis - The purpose of the risk analysis is to determine the impact of individual risks and the likelihood of occurrence. The process of risk evaluation will result in a prioritised list of treatable risks for further consideration.

The result of the risk analysis process will produce a risk profile which gives a significance rating to each risk and provides a tool for prioritising risk treatment efforts.

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This ranks each identified risk so as to give a view of its relative importance. All risks identified will be assessed according to the Risk Matrix and include proposed risk mitigation responses.

Risk Evaluation - The risk evaluation process will document the risks that require mitigation plans and document which resource will be responsible for the mitigation plans.

Risk Treatment - Risk evaluation provides a list of risks requiring treatment. Risk treatment involves identification of the range of options for treating risks, assessing these options and the preparation and implementation of treatment plans.

All risks other than those which are negligible or insignificant should be subject to further analysis and evaluation. Risks which are classified as significant or unacceptable should be further reviewed to determine the best course of action to mitigate the risk, thus reducing the consequence to an acceptable level.

Treatment should follow the hierarchy of controls given below:

Eliminate
Minimise or substitute
Engineering design controls
Isolate
Administrative Controls
Personal Protective Equipment

Monitoring and Review - Monthly risk meetings will be held to capture new risks and to monitor risk mitigation strategies.

17.4.3. Risk Identification and Mitigation

A Hazop 1 formal qualitative Safety and Health risk assessment was completed during the Study. 82 actions arising from the Hazop study were transferred and actioned via the project need list/action log. 40 actions were closed out during the PFS and 42 actions have been retained and will be addressed in the subsequent DFS and execution phases.

A summary of the fifteen highest initial risks with proposed mitigations are listed below:

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Table 113 Project Risk Register

				Inherent Risk			Mitigated Risk
				Likelihood (%)	Exposure		Likelihood after mitigation (%)	Exposure
Risk ID	Risk Category	Risk Description	Risk Realisation When?			Mitigation Plan
6	Infrastructure	Power supply agreement and reticulation to the mine site needs cannot be achieved. Alternative power supply arrangements are required.	DFS	90%	3.60	Combination of Own generation and Grid Power	25%	0.75
8	Site Wide	Ground Conditions Impact on earthworks, foundations, and road construction quantities	DFS	90%	3.60	Geotechnical studies required for FS study and execution.	10%	0.20
14	Process	Ore variability affecting flotation performance	PFS	90%	3.60	Ore Sorting included in Process Design	25%	0.75
25	Tailings	Tailing characterization testwork (Geotech and Geochem) does not match design assumptions. This can impact the rate of rise, final height or liner requirement.	DFS	70%	3.60	Schedule additional geotechnical tests inclusive of critical state analysis and shear tests during the FS. Also complete additional column kinetic tests and radionuclide tests to update the liner system requirement.	10%	0.20
1	Environment	Ionizing radiation of infrastructure due to contamination during LOM.	Closure	75%	3.00	Waste management and closure plan	25%	0.50
13	Process	Sulphuric Acid quantity required neutralization and cost	PFS	75%	3.00	Process design updated to include an acid recovery system to reduce the import of Sulphuric acid	25%	0.75
23	Environment	Radon build up in confirmed spaces	Operation	75%	3.00	Ventilation of Confined spaces. SOP for confirmed spaces and monitoring. Radiation Management plan draft in place	25%	0.50

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				Inherent Risk			Mitigated Risk
				Likelihood (%)	Exposure	Likelihood after mitigation (%)	Exposure
Risk ID	Risk Category	Risk Description	Risk Realisation When?						Mitigation Plan
11	Process	Reagent pricing - Pricing included in the project model affects project value	DFS	75%	2.25	Reagent pricing updated during value engineering	25%	0.50
12	Process	Process reliance on expensive specialist reagents - Florea	PFS	75%	2.25	Investigate and Test alternative suppliers/reagents	50%	1.00
16	Site Wide	Availability of construction power and necessary site support services - water	Execution	75%	2.25	Consider on site boreholes for construction. Contractors to provide their own power generators	10%	0.20
18	Process	Buildup of impurities in the circuits affecting product purity and creating possible processing problems	PFS	75%	2.25	Bleed and water treatment to manage contaminants.	25%	0.50
29	Process	Process circuits adversely affected by buildup of contaminants in return water	PFS	75%	2.25	Bleed and water treatment to manage contaminants.	25%	0.50
4	Environment	Radiation contamination of process equipment cannot be adequately cleaned prior to removal from site. Interim storage is required and disposal of waste on closure is still required.	Operation	50%	2.00	Cat 3 Storage and washing bay, Waste Management and closure plan	25%	0.75
31	Environment	Safety during transportation of reagent to site and product from site	Operation	50%	2.00	SLA with transporters should include that the supplier is responsible for safety during transportation – confirm point of transfer of liability.	25%	0.75

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				Inherent Risk			Mitigated Risk
				Likelihood (%)	Exposure	Likelihood after mitigation (%)	Exposure
Risk ID	Risk Category	Risk Description	Risk Realisation When?						Mitigation Plan
10	Mining	Animal / human interference with trucks on the mine haul roads causes Accidents - especially at night	Operation	50%	1.50	Consider infrared cameras on mobile equipment	25%	0.50

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17.5. Recommendations

The flowsheet as presented has been demonstrated to recover REEs effectively from the Lofdal ore. This flowsheet incorporates several simplifications and improvements to the flowsheet that was presented previously, and metallurgical testing is continuing to further improve the flowsheet.

In particular the following chief risks and opportunities can be addressed in future phases of the project:

Investigate opportunities to reduce the reliance on expensive specialist flotation reagents
Optimise process conditions for the hydrometallurgical circuit based on ongoing testwork using flotation concentrates produced from the upgraded sorted low grade ore samples, including:
Optimise acid bake conditions with respect to REE recovery and impurity control
Refine the acid recovery circuit including bleed requirements

The flowsheet should be updated in the next phase to reflect the optimised conditions when these are available, incorporate the improved REE precipitation circuit and optimise impurity rejection.

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18. Project Infrastructure

The mine, mill processing plant and major mine site related infrastructures will be located at the Lofdal mine site approximately 30 km west of the town of Khorixas are illustrated in Figure 174. The Project infrastructure is designed to support an operation with two (2) open pit mines supplying a 3,010,000 tpa ROM feed to the processing plant, operating on a 365 days per year. It has been developed for the most economical operation at this production rate and will require further expansion and development for any increases in throughput. The overall site layout showing the location of the open pits, processing plant and the waste management site is illustrated in Figure 175.

18.1. Summary

The infrastructure required for the Lofdal REE Project will include:

Mill complex;
Site development and access;
Overall water management plan;
A tailings storage facility (TSF) and associated water management structures;
Electrical site reticulation and generated power;
Warehouse, offices, facilities, and other services.

The proposed layout of the Lofdal Project site is illustrated on Figure 175.

Figure 174 Lofdal Project Site Location

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Figure 175 Lofdal Project Site Layout

A Site Layout scale plan has been produced for the project and represents the overall area where the mining project is to be constructed. The drawing shows the haul roads, A4 open pit, solar PV plant, processing plant with offices, maintenance workshops, and related services.

The mill complex site is located more than 500 m southwest of A4 pit. It consists of the processing plant as well as associated offices, warehouses, maintenance workshops, emergency and first aid services and communal areas like a change house and lunchrooms. The layout of the mill facilities has been optimized to take advantage of topography, reduce earthworks and to position services in proximity to their points of use.

The entire processing plant will cover an area of approximately 153,000 m² and includes the crushing circuits, ore sorting, milling, flotation, acid bake, hydrometallurgical processing, fuel storage and acid storage.

18.2. Bulk Water Supply

18.2.1. Water Demand

All water supply infrastructure concept / basic designs were based on an annual average demand of 1.5 Mm³/annum or average annual daily demand (AADD) of 4 061 m³/day. The water supply infrastructure was designed assuming a 20-hour pumping day, i.e. on a flow rate of 203 m³/h. It was also taken into consideration that the total maximum flow rate the borehole scheme could provide in the future is given as 220 m³/h.

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18.2.2. General

The basis of design is that the water demand of the project will be supplied from 8 boreholes situated north of Fransfontein and pumped to one collector / terminal reservoir with a capacity of 302 m³ situated next to the C35 gravel road.

Figure 176 Assumed Borehole Locations and Pipeline Route

The groundwater supply scheme is based on the geohydrological investigation completed by SLR Consulting. Borehole yield data were sourced from this study, which indicates a total sustainable yield of approximately 220 m³/h across the proposed production boreholes.

Based on the geohydrological assessment and subsequent discussions with SLR, the primary groundwater resource is expected to be derived from the Fransfontein aquifer. The prospective production area, as identified by SLR, is delineated in red in Figure 18-3. The assumed sustainable yield has been benchmarked against the performance of existing NamWater production boreholes in the Fransfontein area.

The geohydrological data formed the basis for detailed borehole modelling and the design of associated infrastructure, including pumping equipment, electrical supply and control systems, rising mains, pipeline routes, and ancillary civil works. Water from the production boreholes will be conveyed to a central collector reservoir located in close proximity to the eight proposed borehole sites.

From the collector reservoir, raw water will flow by gravity through an approximately 45.1 km-long underground oPVC pipeline (OD 315/250 mm) to the terminal reservoir at the mine site. The selected pipeline route and elevation profile eliminate the need for intermediate booster pumping stations.

18.2.3. Boreholes

The borehole infrastructure was modelled on EPANET (Water modelling software) which enabled sizing of all the required infrastructure such as pumping equipment, electrical equipment, piping infrastructure, pipeline routes and civil works required, etc.

All boreholes will be equipped with a stainless steel multistage submersible pump design to deliver at least the assumed yield of the specific borehole at the desired total head. The preferred borehole piping is Boreline flexible dropper pipe (hose). All instrumentation and electrical equipment will be housed in a separate small building at each borehole to protect them. The borehole installations will be securely fenced. The entire borehole installation will comply to NamWater standards to allow infrastructure to be taken over by NamWater to

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operate, if required. The borehole pipeline network will be installed below ground with the piping material of construction either being uPVC or oPVC, depending on the size and application.

18.2.4. Transfer Pipeline

The preliminary pipeline route will be approximately 45.1 km from the collector reservoir to the Lofdal project site. The pipe route will follow the district roads where possible, as indicated. This was done to avoid registering a new servitude for the pipeline and to have an as small as possible environmental impact. Where the pipeline will follow the district roads and be installed in this road servitude the applicable authority should provide their approval. In this case, this is the Roads Authority of Namibia. Further coordination should be completed in the next stage of the project and when the final route is known.

A preliminary hydraulic model generated in EPANET with flows and pressures is shown in Figure 177.

Figure 177 EPANET Model

The recommendation is for a 45.1 km OD 250/315 mm, PN 12.5 / 16 / 20 oPVC gravity line from the borehole Collector Reservoir to Site. oPVC is also the preferred piping material of NamWater for this pressure class and pipe diameter.

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18.2.5. Pipeline Appurtenances

The following appurtenances will be allowed for the pipelines.

18.2.5.1. Inline Isolating Valve Installations

The design should align with NamWater’s design requirements to install inline isolating valves every 4 to 5 km along the pipeline. The valves would be underground and simply marked with a concrete block on the surface or could be installed in a covered manhole for protection.

18.2.5.2. Scour Installations

The number of scour installations between two inline isolating valves depends on the topography and the ability to completely drain the respective section of pipeline. The final positions will be determined at detail design phase when the actual survey of the pipe route has been completed. Each scour valve would be installed in a cage with the discharge pipework facing to the outside.

18.2.5.3. Air Valve Installations

The purpose of air valves is to release air, break vacuums and alleviate surges. Based on the experience on other pipeline projects, the distance between air valves is not to exceed 500 m. The final positions and sizes of the air valve installations can only be determined with the design of the vertical alignment during the detailed design phase. Each air valve would be installed in a cage.

18.2.5.4. Pipeline Markers

It is recommended to mark each point of intersection of the pipeline with a concrete marker block and between the abovementioned manholes and cages where the distance exceeds 250 m.

18.2.6. Water Storage along the Pipeline Route

18.2.6.1. Borehole Collector and Pressure Break Reservoirs

It is proposed to install a 302m³ raw water collector reservoir that will provide approximately 12 hours of storage, if one of the boreholes cannot supply water. This is additional to the 48-hour storage at the mine.

The reservoir / tank allowed for is a circular sectional steel tank with a 7.3 m diameter, consisting of three panels, i.e., 7.2 m high with a capacity of 302 m³. The circular sectional steel tank will be installed on a reinforced concrete ring beam. The tank will be of bolted construction, complete with the channel ring for fixture to the concrete ring beam, the floor membrane of durable vinyl sheeting, a galvanized steel roof with screened ventilators, access manhole and hinged cover, an internal access- and external caged access ladder and water level indicator. The roof will be insect-proof, and the manhole cover will be pad lockable.

The tank shell panels are manufactured from 4.5 mm thick carbon steel sheeting and are hot dip galvanized after manufacture. All pipe connection fittings (inlet-, outlet-, scour-, and overflow pipe connections) will be welded to the respective shell panels during manufacturing and before galvanizing. The roof sheets have a minimum thickness of 2.5 mm.

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All bolts, nuts and washers are hot dip galvanized and of high tensile grade. All sealants are non-tainting, non-toxic and suitable for potable water.

18.2.7. Motor Control Centres

18.2.7.1. General

Motor Control Centres (MCC’s) will be installed to control each borehole pump station. The MCC will be designed according to NamWater standards. The MCC will be a modular type of panel which includes the following:

Main income,
Main isolator;
Surge protection;
Energy demand meter;
Over and under voltage protection; and
Phase failure protection.
Power supply cubicles for major equipment,
Local electrical distribution,
Control section,
Programmable logic controller (PLC);
Human machine interface (HMI); and
Direct current (DC) power supply.

All VSDs will be installed externally of the MCC enclosure as per NamWater’s standard. All VSDs will have harmonic filters and will be of the low harmonic type.

18.2.7.2. Borehole Pump Control

All boreholes will be SCADA controlled via the level in the collector reservoir. Variable speed drives and flow meters will be installed to ensure the pump flows do not exceed the recommended yields of the boreholes. The pumps may also be operated in a manual mode. There will be a low-level protection for the pumps that will override the automatic/manual operation to prevent damage to the pumps if the water in the borehole is below operational levels.

It is crucial that as far as possible, all duty boreholes run simultaneously, to allow sufficient head for the borehole pumps to operate as close as possible to their best efficiency points (BEP). Running a low number of borehole pumps at a time may cause cavitation and cause negative pressure in the main pipeline. The proposed control and operating philosophy should be further investigated during the preliminary and detail design stages of the project.

The pumps will have sufficient protection (mechanically and electrically) to prevent premature failure of the system. Variable speed control will form part of the pump design to regulate the flow from the boreholes.

The borehole flow and totalized flow will be measured. This will include electromagnetic flow measuring devices to provide flow measurements to ensure the water balance can be achieved.

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18.2.7.3. PLC and SCADA

The borehole pump stations will be remotely monitored via a supervisory control and data acquisition (SCADA) system. This would mean that the plant/scheme can be monitored from Windhoek (NamWater), and certain remote-control features will be possible depending on the client's requirements.

A local PLC will control the pumps at each station. The final control philosophy and architecture of the PLC and SCADA system will be completed at final design stage with the approval of the client and NamWater.

The installation will include communication device/s to communicate with a telemetry system via Radio Frequency (RF) or Global System for Mobile communication (GSM). The signal quality should be verified.

The telemetry equipment will be installed in a separate enclosure complying with NamWater standards in a new building.

18.2.8. Electrical Power Supply

Power supply to the borehole pumps stations will be 400 V, 3 phase 50 Hz. All equipment is selected to comply with this specification.

The boreholes will be supplied by 11/0.4 kV transformers rated at 50 kVA (at this stage). 11 kV lines will be extended from the existing NamPower 11 kV line, which is currently feeding the NamWater boreholes at Fransfontein. The line will run close to the boreholes to enable the installation of the transformers close to the boreholes.

Once the final number and positions of the boreholes are known the final route and supply should be agreed with NamPower.

18.2.9. Surge / Lightning Protection, Earthing and Bonding

Surge protection will be provided in all distribution boards, control panels, MCCs, etc., complying in every respect with NamWater's standards.

The lightning protection will also be provided and comply with SANS 10313: Protection against lightning – Physical damage to structures and life hazard. Earthing calculations will only be completed during the detail design stage.

18.2.10. Fire Protection

All borehole pump stations will be provided with fire protection complying with SANS 10400 T. Each MCC room will be equipped with a 5 kg CO₂ fire extinguisher. Due to the low risks of a fire, no additional fire protection or fire detection will be required.

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18.2.11. Pump Station Buildings / Enclosures / Structures

A pump station (brick building) has been allowed with security fencing to restrict access.

18.2.12. Raw Water Storage

18.2.12.1. General

The terminal reservoir will have a minimum 48-hour storage capacity of the AADD, as per normal NamWater standards. Therefore, the bulk terminal reservoir/s should have a minimum capacity of 8 500m³.

The water storage tanks will be constructed from sectional steel tanks with a minimum combined capacity of 8 500m³. Steel reservoir/s are opted for due to their lower capital cost and since the life of mine is estimated at 16 years, concrete reservoirs would not be feasible.

Two Urban 3010 (4 556m³) or equivalent ground level, steel, circular reservoirs were considered for the raw water storage. The tanks are manufactured from Zincalume steel sheets corpus with inner liner founded on a concrete ring beam with a floor liner. The roof is domed and manufactured from Zincalume corrugated steel sheets on hot dipped galvanized steel trusses.

A preliminary design was completed for the concrete foundation ring. The reservoirs will be constructed at a high elevation point, south of the mine site. From here the raw water will be able to gravitate to the required supply points and processes. The reservoir area will be fenced with a security fence. At this stage the envisaged area required is 80 m by 50 m, i.e. 4 000 m².

18.2.12.2. Fire Water Storage

The fire water storage capacity was calculated assuming a high-risk fire with three hydrants operating simultaneously for 4 hours. This requires a total storage capacity of 1 080 m³.

It is recommended to use a dedicated fire water storage tank to ensure sufficient storage is available, even the vent of peak water demand, in the event of a fire emergency. In a peak demand scenario, the water storage may be low and insufficient fire water may be available if a single storage vessel is used.

The tank will be similar to the general raw water storage reservoir and will be constructed with sectional steel members.

The proposed steel reservoir / tank is an Urban 225 (1 224 m³) or equivalent ground level, steel, circular reservoir for the fire water storage. The tank is manufactured from Zincalume steel sheets corpus with inner liner founded on a concrete ring beam with a floor liner. The roof is domed and manufactured from Zincalume corrugated steel sheets on hot dipped galvanized steel trusses.

The consultant has completed a preliminary design for the concrete foundation ring to obtain quantities for cost estimates. The reservoir will be constructed at the same position as the general raw water storage reservoirs to allow for the tank to be installed on an elevated position and enclosed with security fencing. At this stage the envisaged area required is 35 m by 35 m, i.e. 1 225 m².

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18.2.13. Domestic Water Supply

18.2.13.1. General

A new Reverse Osmosis (RO) water treatment plant will draw water from the main / general raw water reservoir and treat the water to comply with the new Namibian Water Act. Once the raw water has been treated it will be disinfected and pumped to six 10 m³ LLDPE tanks installed on one steel elevated tank stand. From here the water will gravitate via the new domestic water reticulation installation to the required domestic supply points, as required.

18.2.13.2. Water Treatment Plant

From the data collected from existing boreholes, indications are that the raw water would require treatment to comply with the Namibian potable water standards. The proposed water treatment plant (WTP) would be a containerized RO plant and would include the following processes:

Raw water pumps
Pre-filtration (pressure filtration / microfiltration)
High pressure RO pumps
Two stage brackish RO membranes and pressure vessels
Blending of raw water
Disinfection
Intermediate storage
Backwash and cleaning in place (CIP) systems
Chemical dosing
All piping, valves, and instrumentation

The capacity of the plant is 30 m³/day and the required raw water supply to the plant is calculated to be 44.4 m³/day. This is a preliminary estimate, and the final raw water requirement will depend on the final water quality.

The preliminary design assumed that the waste and brine from the RO WTP will be integrated into the overall planned wastage of the mine and no evaporation ponds are included at this stage.

18.2.13.3. Potable Water Booster Pump

Once the water has been treated it will be pumped with a booster pump set to the elevated tanks. The booster station will consist of two vertical multistage inline centrifugal type pumps installed in a one duty plus one standby configuration.

The total domestic water demand was calculated to be 30 m³/day, with three shift changes each day and 60 shift workers at a time. Shift workers will be required to shower at the end of each shift. If a flow rate of 6 l/min/person is assumed for a shower, as given in SANS 10252-1, and that each shift worker showers for 10 minutes, the maximum flow rate per hour from the showers would be 60 x 6 x 10 / 1000 = 3.6 m³/h. If it is assumed the rest of the flow stays relative constant over a 20-hour period, the peak hourly flow rate would be 4.7 m³/h and will occur 3 times a day.

The pumps are sized to achieve the hourly peak flow rate and the total head.

The pumps will be level controlled from the elevated tanks.

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18.2.13.4. Potable Water Storage

It is recommended that a minimum 48-hour storage capacity of the AADD, as per NamWater’s standards, be adopted. Therefore, the potable water storage should have a minimum capacity of 60m³. A 15 m high steel structure with six 10 m³ LLDPE storage tanks is proposed at this stage. The tanks will be interconnected and will be provided with isolating valves to allow isolation of any of the six tanks for maintenance purposes.

18.2.13.5. External Potable Water Reticulation

All external water reticulation pipework will be Polyethylene HDPE (PE 100) PN 10, conforming to ISO 4427 or uPVC Class 9 conforming to SANS 966-1.

18.2.13.6. Internal Potable Water Reticulation

All domestic internal hot and cold-water pipework will be copper Class 0, conforming to SANS 460 or PEX-AL-PEX conforming to SANS 21003.

18.3. Bulk Power Supply

18.3.1. Electrical Load

The Lofdal Rare Earths Project is a Greenfields Project without any significant power supply currently available close to the site. The predicted electrical demand load is approximately 12,3 MW (14 MVA) during open pit mining operations. This estimated load is based on the current mill process-mechanical load, mill utility load, tailings management facility load, and auxiliary building load, open pit mine load, ancillary loads, and an allowance for future nominal growth / changes of auxiliary loads over time.

The Namibian power Utility NamPower indicated in December 2024 that they will be able to supply 13,3 MVA (and more) to the site via a +/- 200 km 132 kV line and a local stepdown substation, but that this will be challenging in terms of timelines and costs.

18.3.2. Power Generation

The bulk power supply strategy for the project will be a combination of the below power supply mix options:

Grid Supply via the national energy supply utility, Namibia Power Corporation (Nampower);
Independent Power Producer (IPP) via Renewables and Diesel Generators.

During construction of the mill and mine site, diesel power generation is to be used as the primary power source.

Grid supply will be connected from Gerus substation, between Otjiwarongo and Outjo, via the new +/- 200 km 132 kV line to the new Lofdal 66/11 kV substation. The power infrastructure will be supplemented by an IPP power plant, consisting of a combination of Solar PV, possibly BESS and a Diesel generating station. The entire project will be fed from the main incoming substation at 11 kV, where all integration between the different supplies will be achieved as well.

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Figure 178 Bulk Power Infrastructure

Connecting to the grid in Namibia, depending on the nature of the deep connection infrastructure changes, is known to take around 48 months before a new client can be connected. This is taken from the time a power supply agreement (issued by NamPower) and its contained conditions has been met in full. It was therefore anticipated that grid connected power will have to be supplemented should the project require power before the grid connection has been finalized. The chosen supply mix strategy adds the versatility to the project to have power available in such a case through deploying the IPP timeously.

18.3.3. Transmission Line

The standard NamPower specification is for a single Pelican conductor on a Steel Monopole (Type 259A structure) with an Optical Fibre Ground Wire (OPGW).

Figure 179 Transmission Line Route

The proposed line route is based on the best route in terms of distance and alignment with other infrastructure (e.g. the water pipeline). This route is preliminary and should consequently

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be optimized during further studies to reduce the amount of bend points and towers that may be used through routing the line away from steep elevation changes.

Figure 180 Line Route Elevation

The line route elevation is shown above. From the graph, the elevation changes about 180 m (992 masl to 1070 masl) over the length of the first section of line up to the existing Welwitschia substation area (41.3 km). The valleys and peaks are relatively few, and the line elevation should consequently not provide many construction challenges to a reputable contractor. The rest of the line (from the Welwitschia Substation area to Gerus Substation) will follow the route along the existing 66 kV line from Gerus to Welwitschia substations.

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Figure 181 Proposed Line Route and Bend Points

18.3.4. 132/11 kV Lofdal Substation

As part of the 2024 NamPower enquiry, an estimate was received for a 132/11 kV substation to supply 13,33 MVA of power to the Lofdal Project. CREO has taken the approach of designing a 2 x 20 MVA substation to ensure adequate supply capacity on the transformers under all current and future conditions. In the interim, a 1 x 20 MVA substation should be sufficient until such time as the load for the project is confirmed.

An adequately sized platform of +/- 105 x 105 m, extending from 1000 mm below natural ground level to 500 mm above natural ground level, with a slope of 1:200, constructed with G5-G7 material and installed in layers of 200 mm and compacted to 93% MOD AASHTO is required for the substation. All electrical equipment will be installed on top of this platform.

A sufficiently sized horizontal grid, constructed of 10 mm diameter round copper, which are connected to 15 m copper plated steel, vertical rods are required for the earth mat of the substation and should be installed as part of the platform.

In addition, an access road which is at least 6m wide with a bending radius of 20 m, minimum, should be constructed from the nearest road to the substation level using G5-G7 material and compacted to 93% MOD AASHTO.

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18.3.5. Electrical Works

Figure 182 132/11kV In-take Substation Single Line Diagram (incorporating future second transformer)

18.3.6. IPP Supply

The IPP Power Plant will consist of a PV Generating Plant and Diesel Generators. A Battery Energy Storage System (BESS) may also be included to assist with reduction of diesel usage and for load regulation (e.g., during period of clouds moving over site, for maintenance on generators, etc.).

Although the Modified Single Buyer (MSB) rules of the Electricity Control Board (ECB) of Namibia, limits alternative energy generation to network connected customers to 30% of its consumption under normal conditions, it would be possible to get exemption when insufficient grid power is available.

18.3.7. Solar PV

The optimal sizing of the PV facility in parallel with the grid connection was found to be 14 600 kWp as deducted from the trade-off study. For the 90% renewable option (off-grid), a 47MW Solar PV system would have been required.

The IPP will design, engineer, procure, supply, deliver, construct, install, commission, operate, and maintain a Solar PV Generating plant at Lofdal Rare Earths Project in accordance with NamPower's RE Embedded Generation Specifications.

The IPP will operate and maintain the Solar PV plant for the duration of the contract to supply Solar PV generated power to the mine for the life of mine in terms of the agreement. The final size of the plant can still be optimised during further project development.

The location of the PV generating plant is within the mining license boundary. The IPP will be responsible for all earthworks and site clearance work, civil works and structural works for the safe plant installation and operation. The IPP must include all security infrastructure to ensure the plant is secure, including all roads necessary to maintain the site and equipment on the site.

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The IPP will be fully responsible for obtaining their own EIA related to generation and ensure it will not adversely impact the mine's EIA. The successful bidder will be responsible to timeously provide the necessary information.

It was suggested that the power generated by the plant will be stepped up to 11 kV using a step-up transformer which forms part of the Solar PV generating system substation.

The overhead distribution lines should be designed to 11 kV, according to NamPower specifications, to ensure adequate voltage regulation over the distance from the PV generating facility to the mine in-take substation, from the IPP system at 11 kV, via a step-up transformer and associated control switchgear. Metering and cabling to the mine 11 kV substation will be purchased and installed by the IPP.

The Solar PV generating system will be controlled and operated by a SCADA system, which should interface with the planned processing plant's process control system via fibre cable, if required.

The PV generating plant must be "visible" to Lofdal Rare Earths Project control system via a dedicated network.

The IPP must interface with the project's electrical protection system. The protection schemes for both the project and NamPower shall be compliant with NamPower's grid code for generation systems greater than 5 MW.

When the project load is reduced during operation below the output of the PV plant, power can possibly be fed back to NamPower, or the output of the Solar generating plant is reduced to meet the project's demand. This reduction is achieved by "throttling" the inverter output on the Solar PV system and/or disconnecting some of the Solar PV plant from the electrical network through the existing control systems associated with a Solar PV plant of this magnitude. Should the project be allowed to feed power back to NamPower, this will be an added advantage which could further reduce OPEX in times of low demand (e.g., plant shutdowns for maintenance purposes). This potential needs to be investigated as part of future project development.

A control facility, for limited backup power (after grid connection), will be used for co-generation with diesel generators when NamPower cannot supply power to the mine due to loss of grid. Both PV and Diesel Generator solutions must therefore be supplied by the same IPP, the two solutions are interlinked and cannot be separated.

18.3.8. Diesel Generation

The Diesel Generators should have a capacity of 16 000 kW in order to supply the supply the entire load with power during all contingencies.

The IPP shall design, engineer, procure, supply, deliver, construct, install, commission, operate, and maintain a diesel generation plant that will be used for baseload generation during the initial gap-period (If applicable), while the grid supply is being constructed, that can supply the peak load of the project thereafter.

The plant is situated within the mining license boundary and will be operated and maintained by the IPP. The diesel generators must each be in its own container and the container must be movable.

The IPP will be responsible for all works, Civil, Structural, Installation and Commissioning. This will include all the necessary licenses and permissions required to operate the diesel generation. Although the facility is located within the mining licence boundary the IPP will still be required to complete their own EIA to ensure it will not adversely impact the project's EIA.

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The IPP will make provision for fuel storage to enable 72 hours running at full capacity. The diesel fuel will be supplied by the client and will be monitored by a fuel management system to be agreed between the IPP and the client.

The generators must be sized and configured to be able to start up the 6.5 MW load from a running base load of 2 MW at a power factor of 0.78. The generators will be synchronised to the mine's operating 11 kV system with load sharing capability in an automated system and their status will be visible to the mine SCADA system and connected using fibre cable.

18.3.9. Battery Energy Storage System (BESS)

The purpose of the BESS, as envisaged during the previous study, was to maximise the amount of renewable energy (Solar PV) to be used and to minimise the amount of diesel usage. As NamPower will be supplying the full amount of power required by the mine for the grid connected option, diesel generation would be limited to emergencies only and a BESS is therefore not required in parallel with the grid supply. It is, however, essential for the off-grid supply (to ensure +/-90% renewable energy generation is utilised).

The IPP shall design, engineer, procure, supply, deliver, construct, install, commission, operate, and maintain a BESS system that will be used for Solar optimisation (up to 90%) and diesel consumption reduction. The plant is situated within the mining license boundary and will be operated and maintained by the IPP. The BESS will be containerised, and the container(s) must be movable.

18.4. Electrical Distribution

Loadings of similar sized buildings were used to estimate the electrical loading of each building. These loadings define the design basis of the Internal Electrical Distribution. The total loadings required made it possible to identify the placing of two 315 kVA, 11/0.4 kV minisubs in the vicinity of the buildings. These minisubs will be fed with an 11 kV cable (95 mm²) from a single 11 kV feeder in the 132/11 kV substation yard.

The minisubs in turn feed seven (7) strategically placed kiosks which feed the buildings with power. The feeds from the kiosks are via suitably rated breakers and cables. The kiosks also feed 32 streetlights which are placed around the buildings with the purpose of supplying sufficient light for vehicles and pedestrians to move safely around the buildings during the night.

Nine (9) high mast lights are also required in the plant area to provide sufficient light for the movement of vehicles in this area. These lights will be fed from internally placed kiosks and/or DBs in the plant area.

18.5. Site-Wide Communications

The mine site will employ a site-wide communications system based on a single mode fiber optic backbone. VOIP telephones, intranet/internet access, and control system network connectivity will be integrated into this fiber backbone so that these systems can be accessible anywhere on site. Broadband internet access will be purchased from a satellite internet service provider. The corporate network (intranet) will be isolated from the control system network via a firewalled DMZ (de-militarized zone) network.

The fiber site reticulation design was based on similar sized projects. Both a Main and Emergency Fiber network (with fire retardant fiber) is assumed to be present, and that seven (7) switches will be placed strategically around the mine area.

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18.6. Access Roads

The main access road to the mill complex from the town of Khorixas starts at the D2625 turn-off. The first 20.25-kilometer consists of well-constructed gravel district road in a reasonable condition as observed during a site visit, since it is frequented by tourists and thus maintained to some degree. The road exits Khorixas in a northwestern direction, before turning west and finally south as it approaches the site, see

Figure 183. The last 11.8 km of the route is currently a two-track farm road, that will require gravel road construction for industrial use. The new access road will approach site south of the TSF and A4 pit area to clear areas associated with mining and tailings operations, with a main gate at the site entrance that will be manned 24/7.

Current base layers will be utilized while the wearing course will be redone for the applicable D2625 road. An allowance has been made for low water culvert crossings for access after minor flash floods. The new gravel road will be constructed as per Namibia Roads Authority standards to support two-way traffic of personnel and materials/supplies. This will comprise of clearing, roadbed preparation, cut to fill operation and the addition of base and wearing course layers. The new road is planned to be 10m wide.

Internal haul roads are planned to be constructed for ore and waste transport from the open pits to their designated destinations. Mine haul road and service roads are planned to be constructed to accommodate 30 and 50 nominal tonne trucks carrying ore and waste from the pit to the crusher, and waste to the waste rock area and to the TSF. The waste haul road will connect with the aggregate pit and will also serve as the access to the contractor aggregate primary crusher. The internal roads will require gravel road construction. This will comprise of clearing, roadbed preparation, cut to fill operation and the addition of base and wearing course layers.

Figure 183 Access route via the D2625 from the North

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An alternative route has been identified to serve as a service and construction road, illustrated in

Figure 184. The D2625 East Access route is 29.1-kilometer in total distance and although shorter than the North Access route, it includes a much larger section of two-track farm road, at 21-kilometer. This route would require significantly more capital investment to upgrade and is therefore not a feasible option for a primary access route. It is, however, in proximity to planned bulk power and water supply infrastructure and with it being the shortest route, it is ideal for use during the construction phase of the project and during emergencies whenever the D2625 North Access route is unavailable.

The road exits Khorixas in a northwestern direction, turning west after 8.1 km and proceeds mostly west until it reaches the site. Basic preparation and annual grading of the D2625 East route is recommended.

Figure 184 Access route via the D2625 from the East.

18.7. Plant Buildings

To establish the building requirements, a list of standard facilities was prepared, per section, whilst also considering the workforce that will have to be accommodated in each section. Where departments collaborate daily, the sections were grouped together. Office space was determined by allocating single offices to senior personnel, shared offices for middle level positions and open plan offices for lower-level positions, where applicable. It was decided to have a single change house and washery building at the site entrance with sufficient facilities to accommodate shift changes across all the applicable departments, excluding contractor mining.

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Figure 185 Preliminary buildings layout plan

The design of the buildings and workshops are based on SANS design codes, the National Building regulations and Guidelines for human settlement, planning and design by CSIR and Construction Technology. The design is also influenced by practical and suitable solutions.

For office type buildings, reinforced concrete rafts with prefab units positioned on top were considered. This type of construction is quick, economical, practical and frequently used in similar applications.

The workshops are portal frame type structures, bases and concrete slabs with nominal brickwork where required by function. Additional lean to and cladding type add on requirements will be based on the office type building practice.

18.8. Transportation and Site Vehicles

18.8.1. Site LDVs

To determine the amount of light driving vehicles (LDV's) required on site, personnel count and positions were reviewed, per section, with vehicles allocated as considered necessary. A summary of the allocations can be viewed below.

Table 114 LDV requirement on site.

Site LDV'S Per Section *	Dedicated	Pool
Mining	4	0
Geology	1	1
Safety	1	0
Engineering	6	1

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Site LDV'S Per Section *	Dedicated	Pool
Processing	2	1
Tailings	1	0
Total	18

Excludes provision for contractors.

For the basis of design various models were considered ranging between single cab and double cab as well as 2-wheel drive and 4-wheel drive. Toyota is however the recommended OEM as it is the most reliable and reputable for LDV's in the Namibian mining industry. As part of new purchase agreements, a 4-year or 100,000 km service plan is included.

Table 115 Proposed models for site LDV's.

Make	Range	Model
Toyota	Hilux	Double Cab 2.4 GD6 2x4 SR RB 6MT (A2E)
Toyota	Hilux	Single Cab 2.4 GD6 4X4 SR 6MT (A1L)
Toyota	Hilux	Single Cab 2.4 GD6 2x4 SR 6MT (A1K)
Toyota	Hilux	Double Cab 2.4 GD6 4X4 RAI 6MT (A5P)
Toyota	Land Cruiser Pickup	Double Cab LC79 4.0 Petrol (71G)

18.8.2. Personnel Transportation

For personnel transport to site, it was assumed that the 11 senior positions will utilize private transport. For the middle level positions, three 14-Seaters were proposed travelling mostly on weekdays and for callouts on weekends. The remaining lower-level personnel and shift workers will be transported to and from the site using three 32-Seaters. The frequencies used for the estimate are detailed below.

Table 116 Passenger transport to and from site.

Transport Type	Applicable Personnel	Transport Qty	Frequency	Travel Times
Private	11	10	Weekdays	To mine between 06h00 - 07h00
To town between 16h00 - 18h00
14-Seater	42	3	Mostly weekdays	To mine between 06h00 - 07h00
To town between 16h00 - 18h00
32-Seater	28	1	Mostly weekdays	To mine between 06h00 - 07h00
To town between 16h00 - 18h00
32-Seater	3 x 62 (with 4th shift off)	2	Daily (x3)	Town - mine - town between 06h00 - 08h00
Town - mine - town between 14h00 - 16h00
Town - mine - town between 22h00 - 00h00

The above basis excludes provision for contractors.

The work breakdown structure specifies four employees for every shift position, which indicates a three-shift cycle with a fourth shift on their off-cycle. Regardless of when exactly the shift changes will happen, three shift changes will occur daily during which 62 employees (including front end) will have to be transported to and from site. For the Shift workers, two 32-Seaters were recommended for costing purposes although service providers may propose alternative solutions during service agreement negotiations.

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A high-level trade off evaluation was completed to compare various employee transport options including purchasing of the required transportation, renting of the transportation or hiring a service provider. The results indicated that rental of the required transportation from a rental company whilst appointing the necessary drivers, would prove to be the most feasible.

18.8.3. Material Transport to and from Site

The annual reagent consumption is priced on an ex-site basis, which excludes additional transportation costs to the project location. The estimated annual production of TREO concentrate has been revised to approximately 2,400 metric tons, forming the foundation for determining the outbound concentrate transportation requirements. For inbound logistics, the delivery of spare parts, components, and other ancillary materials is projected to require two truckloads per week.

Table 117 Inbound and outbound transport of reagents, spares, and concentrate.

Goods/Reagents to Site	Links/Week	Ton/Week *
Spares/parts/materials	2	64
Product From Site
Outbound concentrate	1.4	64

Represents the payload.

Considering that road transport is limited to 32 t per load, it was calculated that 104 loads would be required for spares/parts or materials to site and 75 loads for concentrate transport to Walvis Bay Harbour for exports.

Two routes were considered for road transport between Walvis Bay Harbour and the site assuming reagents and spares/parts will be imported, and concentrates exported, via the harbour. The route via Henties Bay was found to be more cost effective since it is the shorter route and since the section between Henties Bay and Khorixas is currently in the process of being sealed.

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Figure 186 Preferred route for inbound and outbound material.

18.9. Plant Fencing

For the design basis, total distances were determined using the preliminary layout for security fencing around the plant perimeter, and standard fencing around specific areas within the plant perimeter. In addition to this, the number and type of access points were also identified for costing purposes.

18.10. Tailings Storage Facility

A PFS design for the Tailings Storage Facility (TSF) was completed in November 2024 to store a total of 40 Mt of tailings at a throughput of 2.16 Million tonnes per annum of tailings to the TSF over a 20 years Life of Mine (LOM). The TSF was designed to 1,005 mamsl and the retaining starter embankment raised downstream to increase stability, with subsequent upstream raises to the end of the LOM. A PFS process optimisation including ore sorting and rejection of low grade ore and strategic stockpiling completed in 2025 (this PFS) reduced the feed to the mill and processing plant to 1.482 Mtpa of slurried thickened tailings pumped to the TSF, for a total storage requirement of 16 Mt of tailings. This reduced storage requirement was used to cost the project but the main earthfill and rockfill embankment size was maintained to keep potential for future expansions.

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18.10.1. TSF Concept Design

The proposed TSF is a valley impoundment with an earth fill and waste rock embankment storing thickened slurry tailings on the eastern side of the site infrastructures layout and main open pit. The site was identified as a preferred location by Namibia Critical Metals for the conceptual design. The TSF comprises an earthfill starter embankment with liner system over the embankment upstream face and basin, and a ring-type deposition system with spigots spaced evenly around the main embankment and sides of the TSF. The TSF development strategy includes an initial downstream raise using placed waste rock from the open pit during the first 2 years of operation, followed by an upstream raising strategy to final elevation. The initial decant water strategy was assumed to be on gravity decanting, incorporating a starter and final decant. A concept study of decant options was conducted for the Prefeasibility study, which identified a siphon system as an adequate alternative to provide more flexibility to maintain the decant intake in the supernatant pool through the TSF life cycle. Provision for both system is included in the PFS.

18.10.2. Storage Capacity Assessment

A total storage capacity of 16.0 million tonnes (11.9 million m³ at 1.35 t/m³) is required at the TSF over the 13-year LOM.

The location of the TSF is maintained as the site selected by NCMI. To accommodate the full tailings volumes over the LOM the crest of the outer wall has to be raised to an approximate elevation of 993 mams, which equate to a total wall height of approximately 28 m. A total freeboard of 1.5 m was assumed to safely store tailings and stormwater to the design inflow requirement. Figure 172 below shows the proposed TSF layout.

Figure 187 Tailings Design Layout

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The filling curve for the TSF are presented in. The tailings storage facility starter wall and final tailings crest levels have been calculated with the stage capacity curves (SCC) developed during the capacity assessment. The starter embankment is sized to contain tailings Rate of Rise (RoR) higher than 2.0 m/year. 2.0 m/year was set as a typical maximum RoR for upstream built TSF particularly over a lined foundation to ensure sufficient time for the tailings to drain and consolidate prior to the next raise. From the stage capacity curves it is determined that it would take approximately 3.5 years of deposition to reach a rate of rise below 2.0 m/year. Above this elevation upstream wall raising with tailings can safely commence and raise up to the final crest elevation of 993 mamsl. At this crest elevation, the facility will have an approximate total height of 28 m.

Figure 188 TSF Filling Curve and Estimate of Starter Wall Requirement

18.10.3. Foundation Characteristics

Geotechnical field tests were carried out in March 2024 to characterise the TSF area. This involved the excavation of 26 test pits and drilling of 4 rotary core boreholes, strategically located to evaluate the TSF starter wall area, the TSF basin area, the zone of influence downstream of the TSF and to identify potential borrow sources. In addition, two of the test pits were positioned near to the proposed waste rock dump to determine the variability across the site

The investigated area is primarily underlain by metamorphic rocks, mainly amphibolite with some gneiss and quarzitic gneiss. Minerals like quartz, feldspars, garnet, and micas are found sporadically within the amphibolite. The hillsides are covered by a thin layer of colluvium and nodular calcrete in some areas.

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Topsoil colour varies, indicating different underlying materials: reddish for shallow weathered amphibolite, brown or grey for amphibolite, and light brown for calcrete. Alluvium is present along riverbeds, with deeper profiles observed in boreholes drilled there compared to slopes.

Highly weathered rock outcrops suggested the underlying rock types would be soft to medium hard. UCS tests confirmed shallow rock consistency ranges from soft to medium hard, with one sample indicating hard rock. Groundwater was found at 3 meters below the surface in one borehole. There is a potential risk of leachate migration from the TSF along riverbeds if unlined, necessitating further hydrogeological investigations.

Laboratory testing was conducted in April 2024, on the tailings sample that was used during the alternative tailings technology study conducted earlier in the study to determine the feasibility of storing the tailings as a filtered stack. The laboratory testing was completed at GHD Limited (GHD)'s Waterloo, Ontario, Canada.

18.10.4. Tailings Material Testing

The results of the test work are summarised below and detailed in KP PFS Design Report.

The tailings sample is predominantly fine and comprises of 88% silt and 12% clay. It is non-plastic and classified as a silt (ML) according to USCS.
The average specific gravity of the tailings sample is 2.7;
Standard Proctor test yielded a MDD and OMC of 1.590 t/m³ and 16.4%, respectively;
Modified Proctor test yielded a MDD and OMC and 1.709 t/m³ and 13.90% respectively;
Undrained peak friction angles ranged between 24° and 28° at different effective stresses, whereas undrained residual friction angles ranged between 33° and 34°;
Drained residual friction angles were measured as 34° and 37°;
The average vertical coefficient of permeability for the mixed tailings sample was measured to be 6.50 × 10⁻⁸ m/s, indicating a moderate permeability.

A geochemical assessment was conducted based on the results of the sample tested at SGS laboratories in Lakefield, Canada. The results of the geochemical assessment are summarized below

The mineralogy of the sample is dominated by albite, a silicate rich mineral. The sample contains other silicate and carbonate minerals in lesser percentages. Both the silicate and carbonate minerals will provide some buffering capacity.
The sample does not contain sulphate or sulphide minerals.
Based on the ABA results, sample was classified as non-potentially acid generating (non-PAG).
Leachate results were screened against the IFC effluent standard for mining and the Namibian effluent limits. EPA guidelines were also used to screen Uranium and Thorium.
pH value is alkaline and within IFC effluent standard and Namibian discharge limits.
Arsenic, which is a major constituent of concern in most mines is below the limits.
All major ions were within guideline limits where guideline values were defined.
Iron (Fe) and aluminium (Al) exceed the applicable limits. This is to be expected as aluminium oxide and ferric oxide are part of the Rare Earth Oxides group.
Uranium concentration (0.000871 mg/L) is well under the maximum contaminated level of 0.03 mg/L as per the EPA guidelines.
Thorium is a naturally occurring radioactive metal existing as Th-232, Th-230 or Th-228 and is solid under normal conditions (meaning that it is unlikely to be found in groundwater). There are also man-made forms which are also radioactive. The

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presence of Th in the leachate is noted and should be verified as well as the potential for radioactivity in the tailings material, particularly as it could be in the dust from the tailings material and presenting a risk to inhalation.

In terms of radioactivity, 1 Bq/l of U 238 = 80.9 ug/l of uranium for alpha radionuclides. At 871 ug/l = 10.7 Bq/l which is above the target water drinking water quality range of 3.8 Bq/l with a cancer risk in the range of < 1:200,000 but significant risk of chemical toxicity. Its not clear how much radiation (also as alpha particles) the Thorium could contribute to the overall radioactivity of the tailings material.
The tailings leachate water will not be released but could be associated with seepage and ponding in the tailings facility, so it is recommended to undertake the testing of the radionuclides of the tailings material to determine the potential radioactivity.

In summary, the geochemistry analysis completed during PFS testing program indicates that a liner isn't required from an acid generation or heavy metal leaching but that the tailings indicate naturally occurring radioactive metals which are not yet fully understood in terms of risks impact to the TSF workers, groundwater and air quality. The foundation characterization would also likely induce high seepage rates to the underground and high losses in the water balance. As such, liner was adopted for the PFS as a precautionary measure, with an opportunity to reduce the lined area to the basal and embankment face areas in the next design stage.

18.10.5. Design Criteria

The design for the TSF is primarily based on the Global Industry Standard on Tailings Management (ICMM; UNEP; PRI, 2020), as well as local and international guidelines for mine waste management design, surface water management design and infrastructure design (Canadian Dam Association - CDA, 2014/2019; Mining Association of Canada (MAC), 2021; Namibia Roads Authority Drainage Manual, 2014a; Namibia Roads Authority Materials Manual, 2014b; the South African National Standards (SANS) on Mine Residue 10286).

The table below summarizes the design criteria used for the PFS design.

Table 118 Design Criteria-Summary

Item	Design Criteria
Topographical Survey	June 2020 topographic survey
Legal Framework	Minerals (Mining and Prospecting) Act of 1992
	Environmental Management Act of 2007
	Global Industry Standard on Tailings Management (2020)
Documentation	Lofdal Rare Earths Project – Tailings Storage Facility Conceptual Design Report – KP, 2022
Specific Gravity (Solids)	2.697
Particle Size Distribution	P80 = 38 - 45 microns (silts and finer)
Tailings solid content by weight	46 % (SGS)
Average settled dry density (Year 1)	1.25 t/m³
Average settled dry density (after Year 1)	1.35 t/m³
Standard Proctor maximum density and optimum water content (OMC)	1 590 kg/m³ at 16.4 % OMC

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Item	Design Criteria
Modified Proctor maximum density and OMC	1 709 kg/m3 at 13.9 % OMC
Average annual throughput to TSF	1.48 Mtpa
Design Life	12 years (including ramp up)
Rate of Rise	2.0 m/year for tailings not impounded behind starter wall (justified by the fine-grained nature of the tailings)
Storage Capacity Required	16 Mt
Tailing’s chemistry	Non-acid generation, non-detectable sulphides and neutralization potential. There is a potential for radioactivity due to presence of radioactive metals.
Slope Stability	The recommended Factor of Safety for slope stability under static loading for construction, operations and transition phases as per CDA, 2019 will be:
Long term (steady state seepage) -Static: 1.5
Pseudo-static: 1.0
Full or partial rapid drawdown - Undrained Peak 1.3
Post-earthquake -Undrained Residual: 1.2
Overall Outer Side Slope	1V: 3.5H
Conceptual Closure Design	The TSF side slopes are to be cladded with 750 mm thick layer of waste rock. There will be no bench drains. The TSF will be capped with 200 mm of topsoil/seedling.
Inflow design flood / hydrologic design criteria	1 in 2,475 years storm event (24 hrs) (to be safely contained and passed through water management system)
Seismic design criteria	1 in 2,475 years seismic event

18.10.6. TSF Components and Geometry

The following are the overall geometry of the TSF

Starter embankment elevation/height: 981 mamsl, 16 m height
Starter embankment volume: Approximately 190 000 m³
TSF final elevation: 993 mamsl.
TSF final height: 28 m.
Overall outer slope: 1 vertical to 3.5 horizontal (1v:3.5h) including benches, waste rock buttress and upstream raises.
Conceptual freeboard requirement for IDF storm event: 1.5 m

Starter embankment: The TSF will consist of a starter wall with crest width of 10 m and crest elevation of 981 mamsl. The height of the starter wall will vary from 0 to 15 m high across the valley. The starter wall will be constructed from overburden earthfill borrowed from its footprint and areas surrounding the TSF with an outer wall slope of 1V:3.5H and an inner wall slope of 1V:2H and placed on shallow bedrock.

Waste Rock Embankment: The TSF embankment will be enlarged with open pit waste rock and constructed downstream of the starter embankment to elevation 986 maml during the initial two years of operation. This will create additional storage and increase stability of the

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overall slope. The waste rock embankment will 20 m wide and 21 m high lift to elevation 986 mamsl with an intermediate bench at 976 mamsl.

Outer Wall: Above the waste rock embankment upstream construction with tailings will entail the following:

Establishment of an appropriate step-in from the previous embankment raise to maintain an overall slope angle of 1V:3.5H. The outer wall will be raised to El. 993 mamsl, equating to a total height of approximately 28 m.
Construction of consecutive machine-built 1 to 1.5 m high tailings lifts around the TSF perimeter.
Filling of the 1 – 1.5 m lifts with tailings via spigots
Paddocks to be constructed at the toe of the embankment to store any runoff material. The material will be used to maintain the outer wall of the embankment
Starter embankment elevation/height: 981 mamsl, 16 m height

Decant System: A concept study was undertaken during this stage of design evaluating the following 4 different decant solutions. The concept study determined that siphon decant system maybe the most feasible option for decanting water from the TSF followed in ranking by the land-based pump system. For the siphon system an intake pipe will be placed at the pool to decant water over the embankment. As the TSF is raised the siphon system will also need to be raised. The siphon inlet point can be relocated as the pool moves. As the TSF is raised at the pool covers a larger area, multiple inlet points can be used to manage the pool size and location.

Return Water Dam: The RWD should have a total storage of 90,000 m³ plus freeboard to store decant water, and water event requirement for the dam classification. It is envisaged to split the RWD into two compartments, for operational purposes and evaporation mitigation.

TSF and RWD Liner: A single HDPE liner with underliner and underdrainage system similar to a “Class C” liner in accordance with GN636 (2013) is included in the PFS design along the starter embankment upstream face and initial TSF deposition area. The liner system is to reduce potential seepage and mitigate long term effect of iron (Fe) and aluminium (Al) which exceed the Namibian effluent discharge limits. The proposed liner system will comprise a 150 mm bedding layer with a 1.5 mm textured HDPE geomembrane placed on top and covered by a protective A6 bidim or similar. The overliner underdrainage system is to improve tailings drainage and consolidation but also increase water recycle to the process facility.

18.10.6.1. Key Quantities

Starter wall earth fill volume: 190 000 m³.
On-going waste rock fill placement for wall raising: 85 000 m³ / year (for 2 years but could be sustained afterwards).
Embankment face and basin HDPE liner surface area: 569 000 m² in year 1, and 757 000 m² over year 1 and year 2 to complete the basin lining along the valley impoundment.

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18.10.7. Dam Safety Classification

18.10.7.1. Dam Breach Analysis

The dam breach assessment (DBA) as not been amended following the updated mining schedule and has been carried out at the initial final elevation of 1,005 mamsl. The scoping level breach scenario looked primarily to a flood induced breach (erosion and overtopping). Given that the tonnages reporting to the TSF have reduced it is expected that the breach consequences will be similar or less severe than described in terms of depth and velocity of tailings. It is anticipated that the zone inundated will be similar, carried by the storm event. The DBA will be revised accordingly in the next stage of design.

Only one failure scenario was evaluated for the scoping level assessment. This scenario was modelled as a rainy-day failure event, based on the final embankment crest elevation of 1005 masl. The failure scenario that was modelled assumed that an embankment slope failure would occur at the maximum embankment height section, leading to loss of freeboard and overtopping.

No permanent population at risk resides within the expected area of inundation, but mine workers are expected to be located within the Open Pit A4 area. Unmitigated, nearly all of the breach effluent would report to the pit, impacting both workers and the mine workings. If sufficient mitigation measures are implemented such as large diversion berms, then the flood wave the majority of the flood wave will divert towards the Huab conservancy instead. It remains important to note that regardless an emergency preparedness and response plan will be required for the open pit area and potentially affected downstream areas.

The modelling results indicates that the dam breach flood wave would result in incremental flooding extent impacts to a point approximately ~10 km downstream of the TSF. below illustrates the results of the modelled scenario.

The contamination of the Kuab River tributary and surrounding areas may result in an environmental impact that may last for years. The quantification of the flow distance and concentration of chemical pollutants was not included in this study.

Figure 189 Inundation zone for rainy day scenario

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18.10.7.2. Dam Safety Classification – Consequence of Failure

The safety/consequence classification is assigned for each individual consequence category outlined in Appendix C of the KP Tailings Storage Facility PFS Design Report. At PFS level, a dam consequence classification of "High" is recommended for the Lofdal TSF in terms of the GISTM Classification Matrix. The classification can be mainly attributed to the possible environmental impact of the breached tailings, associated remediation time and costs, and the anticipated disruption to business due to tailings deposition in the main pit.

18.10.8. Preliminary Seepage and Stability Analysis

The TSF global slope stability analysis presented is based on the initially design final elevation of 1 005 mamsl for 40 Mt storage of tailings. Since the revised final height of the TSF to 993 mamsl does not introduce conditions that would adversely slope stability or facility safety. A comprehensive stability analysis will be carried out in the definitive feasibility stage to evaluate the stability of a reduced TSF height, depending on the FS mining plan. This slope stability model however confirms that the TSF could be raised to elevation 1 005 mamsl if the LOM is extended.

The TSF embankment stability was evaluated assuming that tailings above the phreatic surface are drained, and tailings below the phreatic surface are undrained, this for both peak shear strength conditions and post-liquefaction loading conditions. Figures below show the geotechnical parameters for the material properties used in the seepage / slope stability analysis and shows a typical section used for the analysis. It assumes a lined embankment and basin acting as a weak layer, earth fill starter embankment and downstream waste rock zone.

Figure 190 Cross Section of the starter wall

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Figure 191 Cross Section and Material of the Final Landform (tailings at 1005 mamsl)

Table 119 Seepage/Stability Analysis – Assumed Geotechnical Parameters

Material	Permeability (k) m/s	Saturated unit weight (kN/m³)	Strength Parameters			Source
			c' (kPa)	φ' (°)	shear strength ratio (s₀/σᵥ')
Tailings Material (drained)	6.50E-08	18	0	34	-	(KP, 2024)
Tailings material (undrained peak)	6.50E-08	18	0	-	0.27	(KP, 2024)
Tailings Material (Undrained residual)	6.50E-08	18	0	-	0.19	(KP, 2024)
Compacted fill	1.0E-07	20	0	35	-	Estimated
Waste rock	1.0E-03	21	0	38	-	Estimated
Base Liner (peak)	5.0E-11	2	0	15	-	Estimated
Base Liner (residual)	5.0E-11	2	0	10	-	Estimated

NOTES:
1. Φ', INDICATES EFFECTIVE FRICTION ANGLE
2. C', INDICATES EFFECTIVE COHESION

The results of the slope stability indicate that the proposed design of the TSF starter wall and final landform is safe against slip failures. The obtained FoS against global and local failures for all loading conditions analysed were all satisfactory with respect to static long-term, seismic, short term and undrained conditions as presented in Table 120. Therefore, the proposed TSF is stable and complies with CDA (2019) guidelines. A sensitivity analysis was carried out on drains working and drains not working within the beach to assess the phreatic

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surface level. The seepage analysis indicated that the stability of the TSF is greatly affected by the level of the phreatic surface and the FoS is sensitive to the functionality of the drains. When drains are not working, the phreatic surface rises slightly and the FoS decreases. Tailings used have very fine particles and it is anticipated that in the case were drains are blocked, the stability of the starter wall is adequate, however in the long-term, the stability of the final

Table 120 Slope Stability Analysis Results

Loading Conditions	Loading Conditions
	Static	Pseudo-static	Undrained Peak	Undrained residual (post-liquefaction)
Minimum FoS Criteria	1.5	1.1	1.3	1.2
Starter Wall with Drains
Global failure - Factor of safety through tailings	3.42	2.67	3.53	3.65
Local failure - Factor of Safety through waste rock	2.52	2.12	2.53	2.53
Starter Wall with no Drains
Global failure - Factor of safety through tailings	3.22	2.65	3.5	3.42
Local failure - Factor of Safety through waste rock	2.51	2.13	2.53	2.53
Final Landform with Drains
Global failure - Factor of safety through tailings	3.21	1.69	2.31	1.31
Global - Factor of safety through tailings & buttress (upstream side)	2.98	2.32	2.93	2.88
Local failure - Factor of Safety through buttress	2.51	2.14	2.53	2.51
Final Landform with no Drains
Global failure - Factor of safety through tailings	3.11	1.57	2.20	1.26
Global - Factor of safety through tailings & buttress (upstream side)	3.16	2.24	2.97	2.77
Local failure - Factor of Safety through buttress	2.44	2.14	2.55	2.51

18.10.9. TSF Development and Operational Philosophy

At the starter embankment and waste rock lift, the slurry tailings will be deposited by spigotting as this a cost-effective method of deposition for the fine tailings. The tailings will be delivered via a main delivery line with smaller outlets (spigots) located along the crest of the embankment. The tailings delivery system will include a combination of a gravitational discharge and use of a centrifugal slurry pumps located after the thickeners. The pumps will be used to initiate and maintain gravitational flow of the tailings to the TSF, while the centrifugal pumps further aid in keeping the solids in a fluidized state, thereby reducing the friction head and risk of clogging in the pipes. The slurry will be delivered to the TSF by a 220 mm / 250 mm diameter HDPE pipe and subsequently distributed around the TSF by a 160 mm diameter HDPE pipeline. At any time, the main tailings delivery pipeline will tie to a spigotted pipeline system on the crest of the outer wall along the perimeter of the TSF. From the waste rock lift when transitioning to the upstream raising method, the spigot pipes will be sequentially relocated to the next elevation as the dam develops in 1.0 – 1.5 m lifts

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At this stage a siphon system is recommended for the management of water on the TSF. An intake pipe will be placed at the pool to decant water over the embankment. As the TSF is raised the siphon will also need to be raised. The intake point of the siphon system can be relocated at the pool moves. As the TSF develops, multiple intake points can be constructed to manage the pool size and location on the TSF.

18.10.10. TSF Water Management

18.10.10.1. Climate

The Lofdal Rare Earths Project is located approximately 25 km west of the Khorixas Town, in the Kunene Region of Namibia. The climate in Khorixas can be described as semi-arid to arid, with average summer temperature reaching into 40 degrees Celsius (°C) and winter temperature touching 0°C. The region is a summer rainfall region with the highest temperature and rainfall depths are recorded from December to May.

The climate in Namibia is highly variable, with extreme drought periods and rainfall events (MET, 2011). Climate change models indicate that Namibia, especially the eastern and southern parts are adversely affected by rising temperatures and the consequences thereof (WBG, 2021).

18.10.10.2. Rainfall and Evaporation

Daily rainfall data was received from the Namibian Meteorological (NMET) services for the Khorixas Station for a record period of 56 years from the year 1955 until 2008 (NMET, 2021). Monthly and daily synthetic rainfall data was also obtained from the Climate Research Unit (CRU) and KNMI database (CRU, 2021) (KNMI, 2021), the records did not correlate well with the actual observed dataset, and it was therefore decided to only use the NMET Khorixas Station record.

The monthly rainfall distribution as obtained from the NMET and the Pan evaporation from the Namibia Department of Water Affairs is summarised in Table 121. The mean annual Pan evaporation is 2850 mm (DWA, 1988) and the mean annual precipitation of 223 mm (NMET, 2021). The months with the highest evaporation are December and January and rainfall are February and March.

Table 121 Mean Monthly Rainfall and Evaporation

Monthly Average (mm)	Oct	Nov	Dec	Jan	Feb	Mar	Apr	May	Jun	Jul	Aug	Sep	Total
Rainfall	3	14	16	46	61	60	20	1	0	0	0	2	223
Evaporation	228	342	399	456	428	285	228	143	57	57	86	143	2850

18.10.10.3. Extreme Rainfall Estimation

Daily resolution rainfall data was used to determine the statistical frequency distribution of events. The General-Extreme Value (GEV) resulted in the best fit between the actual observed ranked annual maximum rainfall depths and the various distributions, such as Log-Normal, Log-Pearson and Extreme Value. A Weibull plotting position was used to successfully match the observed data to the distribution. The rainfall depths for different return periods are summarised in Table 122 (SANRAL, 2013) (NRA, 2014).

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Table 122 24-Hour Duration Extreme Rainfall Depths Estimates

Return Periods (years)	5	20	100	500	1000	2000	5000	10000	PMP ¹ (100 000)
Rainfall Depths (mm)	58	90	130	176	197	220	251	277	377

Notes:
1. World Meteorological Organisation (WMO) probable maximum precipitation method could not be applied, due to lack of sub-daily (< 24-hour) rainfall data.
2. 1:100 000-year Return Period rainfall depth derived with the statistical method used as PMP Equivalent.

18.10.10.4. TSF Stormwater Management

A high-level stormwater management assessment was conducted on the surrounding catchments, and they were found to be small in size, as they form the upper reaches of the Huab Catchment. It was found that no storm water diversion system is required upstream of the TSF. It is unlikely that the runoff generated in the catchment adjacent to the toe of the TSF will have an impact on the TSF, as there is sufficient distance between the main watercourse, the RWD and the toe of the TSF.

18.10.10.5. Water Balance

A water balance model was developed, based on the available NMET record and the mine plan. Deposition plan information was used to establish the governing TSF inputs for the water balance. A monthly deposition rate of approximately 123,500 tonnes per month was used in the calculations based on the plant throughput of 1.48 Mtpa. The decant system was sized to handle a PMP of 377 mm over the direct and external catchment (287 ha) to provide assurance that the upstream raised facility could handle extreme rainfalls. A schematic is shown in Figure 192, indicating all the components comprising the water balance.

Figure 192 Simplified Inflows and Outflows Schematic

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Figure 193 Inflow, outflow and decant average monthly volumes

Figure 194 Inflow, outflow and decant wet monthly volumes

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Figure 195 Inflow, outflow and decant dry monthly volumes

The average inflow, outflows and available water for return are presented in the following graph.

Figure 196 Inflow, outflow and decant summary monthly volumes

The PFS TSF water balance shows that overall the water pond is expected to reduce over time with a risk of water shortage during the dry season towards the end of LOM, amplified in dry conditions. This however should be refined in the next design stage to increase sensitivity and probabilistic analysis. It is anticipated that overall the return water will average approximately 46% under average conditions.

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18.11. Plant Sewerage Treatment & Distribution

The total sewer water flow is assumed to be as high as 90% of the total domestic water demand since the sewer flow would originate mostly from offices and ablution facilities. The total domestic water demand was calculated to be 30 m³/day, therefore, the average daily inflow to the sewage treatment plant was calculated to be 27 m³/day. Three shift changes will occur each day with 60 shift workers required to shower at the end of their shifts. If a flow rate of 6 l/min/person is assumed per shower, as given in SANS 10252-1, and it is assumed that each shift worker showers for 10 minutes, the maximum flow rate per hour from the showers would be 60 x 6 x 10 x 0.9 / 1000 = 3.24 m³/h. If it is assumed the rest of flow stays relative constant over a 20-hour period, the peak hourly flow rate would be 4.1 m³/h. This peak will occur 3 times a day.

The sewage inflow quality to the treatment plant was assumed to be as follows:

Chemical Oxygen Demand (COD): 1 000 mg/l
Biological Oxygen Demand (BOD): 500 mg/l
Ammonium (NH₄-N): 60 mg/l

The proposed wastewater treatment plant will consist of the following major process:

Screening (Inlet Works)
Anaerobic Digestion (Septic Tank / Settler)
Anoxic Conditions (Trickling Filter Basin)
Trickling Filter Recycle Sump
Aerobic (Trickling Filter)
Clarification (Secondary Settler)
Disinfection
Polishing Filtration

The effluent discharged from the wastewater treatment plant will comply in every respect with The Department of Water Affairs' Water Resources Management Regulations: Government Notices No 269 of 2023, Water Resources Management Regulations, of the Water Resources Management Act and the effluent could be used for irrigation.

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18.12. Design Parameters

The parameters listed in Table 123 were used as the design basis for the infrastructure development study.

Table 123 Design Criteria Basis
| Design Criteria | Qty | Unit |
| --- | --- | --- |
| Life of Mine | 12 | years |
| Operation | 359 | days per annum |
| | 24/7 | |
| Run of Mine Production | 3 010 000 | tonne per annum |
| Project Electrical Consumption | 94 361 597 | kWh per annum |
| Project Electrical Demand | 12.3 | MW |
| Project Electrical Demand | 13.7 | MVA |
| Water Demand | 1 482 192 | m³ per annum |
| Construction Staff | 300-340 | personnel |
| Operational Staff | 226 | personnel |
| Total Plant Personnel | 243 | personnel |
| Total Mining Contractors | 93-242 | personnel |
| Total Front-End Personnel | 126 | personnel |
| Total Mine Personnel (excluding Mining contractors and Plant) | 38 | personnel |

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19. Market Studies and Contracts

19.1. Introduction

Lofdal is a heavy rare earths project with Dysprosium (Dy), Terbium (Tb) and Yttrium (Y) being the main economic drivers, while the typical magnet light rare earths elements, Neodymium (Nd) and Praseodymium (Pr) will add appreciable value.

Over the last 3 years the rare earth sector has evolved significantly. Demand for magnet rare earths has remained strong, new policy measures have been introduced in major economies, and several peer projects have advanced to Prefeasibility or Definitive Feasibility Study stage.

In addition, yttrium has emerged as a critical pinch point in global supply chains, particularly for aerospace, semiconductors and high-temperature energy applications, following new Chinese export controls in 2025.

This update integrates those developments and benchmarks Lofdal against comparable projects. It focuses on the main economic drivers for Lofdal's basket:

Dy and Tb as critical high-value dopants in high-temperature NdFeB magnets;
Nd and Pr in permanent magnets – NdFeB and SmCo magnets for electric vehicle (EV) traction motors, wind turbines, industrial motors, appliances, electronics, robotics, defence and aerospace;
Yttrium (Y) as a key element in ceramic/oxide high-temperature coatings in turbines and jet engines, specialty alloys and Y-based coatings and insulators used in advanced chip fabrication equipment.

19.2. Overview of Rare Earth Elements

Rare earth elements ("REEs") comprise the lanthanide series plus yttrium and scandium. They are typically grouped into light rare earth elements (LREEs) and heavy rare earth elements (HREEs) based on atomic number and electron configuration. LREEs (e.g., La, Ce, Pr, Nd) are more abundant in the Earth's crust and dominate global tonnage, while HREEs (e.g., Dy, Tb, Ho, Er, Tm, Yb, Lu, plus Y) are scarcer and account for a relatively small share of total rare earth oxide ("TREO") output.

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19.3. Demand for Magnet Rare Earths and Yttrium

19.3.1. Magnet Demand Trends

Global demand for NdFeB magnets has continued to grow strongly since 2020:

EV traction motors, e-mobility (two/three-wheelers, e-bikes, scooters), and wind turbines are the largest structural growth drivers;
Consumer appliances, industrial automation and robotics provide broad-based additional demand;
Independent market studies and OEM disclosures indicate high single-digit to low double-digit CAGR in NdFeB magnet volumes over the 2020–2030 period, consistent with or above the previous forecasts from organizations such as Adamas Intelligence, CRU International and Benchmark Minerals.

Dysprosium and Terbium remain essential dopants for high-temperature NdFeB grades used in EVs, wind turbines and defence applications. While per-magnet Dy loadings have declined through thrifting and improved grain boundary diffusion, the overall demand for Dy/Tb is expected to increase over the life of the Project due to growth in magnet volumes.

19.3.2. Yttrium Demand and Strategic Importance

Yttrium demand is smaller in tonnage than NdPr or Dy/Tb, but recent developments show that it has become strategically critical:

In aerospace, yttrium oxide is used in thermal-barrier coatings for advanced jet engines; U.S. industry regards yttrium as essential for the world's most advanced engines.
In semiconductors, yttrium is used as a protective coating and insulator in critical equipment. Industry sources quoted by Reuters described current yttrium availability issues as a “9 out of 10” in severity, highlighting the risk to production times, costs and equipment efficiency.
In energy, yttrium-containing coatings protect gas turbine blades in power plants from high temperatures.

Given these applications, even modest supply disruptions can have disproportionate economic impact across aerospace, energy and semiconductor sectors.

Table 124 Rare Earth Applications and End-Uses fall into one of Eight End-Use Categories

End-Use Category	Description
Battery Alloys (La, Ce, Pr, Nd)	Rare earth elements are used to produce anode materials for nickel-metal hydride (“NiMH”) batteries. NiMH batteries are used in hybrid electric vehicles, consumer electronics, cordless shavers, cordless powertools, baby monitors and other applications of rechargeable batteries.
Catalysts (La, Ce)	Rare earth elements, such as cerium and lanthanum, are used in catalytic converters of gasoline- and diesel-powered vehicles, as well as fuel cracking catalysts and additives used by oil refiners to break down crude oil into lighter distillates, such as gasoline, diesel, kerosene and more.

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End-Use Category	Description
Ceramics, Pigments and Glazes (La, Ce, Pr, Nd, Y)	Rare earth elements are used to produce decorative ceramics, functional ceramics, structural ceramics, bio ceramics and many other types of ceramics used in everything from jet engine coatings to ceramic cutting tools, dental crowns, ceramic capacitors, ceramic tiles, and more.
Glass Polishing Powders and Additives (Ce, La, Er, Gd, Y)	Rare earth elements, such as cerium, are used to polish optical glass, hard disk drive platters, LCD display screens and gemstones, among a long list of applications. Cerium is also used as an additive in UV-filtering glass and container glass, whereas lanthanum, yttrium and gadolinium are used to produce high quality optical glass used in camera lenses, microscopes and telescopes.
Metallurgy and Alloys (La, Ce, Ho, Gd, Y)	Rare earth mischmetal (a mixture of light REE metals) is used during production of some types of steel, as well as ductile iron making. Rare earth elements are also used to produce a variety of different alloys, such as ferro-cerium, ferro-holmium, ferro-gadolinium and a growing list of others.
Permanent Magnets (Nd, Pr, Dy, Tb, Sm)	Rare earth elements are used to produce high-strength permanent magnets that have enabled the production of ubiquitous gadgets and electronics, such as mobile phones and laptops, as well as power dense energy-efficient electric motors and generators used in electric vehicles, wind turbines, energy efficient appliances and hundreds of other applications.
Phosphors (Ce, La, Y, Tb, Eu)	Rare earth elements are used in phosphors for energy efficient lamps, display screens and avionics, and are added to fiat currency in some nations as an anti-counterfeit measure.
Other (La, Ce, Nd, Dy, Tb, Gd, Lu, Tm)	Aside from the above-described end uses and categories, rare earth elements are used in a long list of other end uses and applications, including many in defense, medicine, aerospace, agriculture, high-tech and chemical industries.

19.3.3. Global REE Demand by 2040

From 2024, CRU expects the REO market to grow at a CAGR of almost 2.4% to 2040, driven primarily by growing demand for permanent magnets (NdFeB) used in EVs, wind turbines, and consumer electronics (see the following two pages for more details on this trend).

Currently, permanent magnets account for more than one-third of the total demand for rare earth oxides (REOs), CRU expects that by 2040 this segment will grow to ~51% of the global REE market.

While demand for polishing powders remains resilient and is forecast to continue growing at a CAGR of 4% annually between 2024 and 2040 – cerium will continue to be in excess supply. Cerium oxide is widely used as a polishing for various materials, particularly glass and optical components.

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Figure 197 Global REO demand by application (in kt REO, source CRU)

Historically, REO demand for end-use in magnets accounted for a modest share of REO demand (around 20% in 2018).

Growth in demand for REOs in magnets has already been outstripping other demand sources due to the use of these magnets in energy transition related industries (primarily in electric vehicles and wind turbines) and this trend is set to continue.

By 2040, BEVs and wind turbines will account for ~26% and ~20% of total NdPr demand up from ~24% and ~14% in 2024.

Magnet REO demand will increase by 87% between 2024 and 2040, whereas non-magnet REO demand will increase by just 9% over the same period.

The point of initial demand for oxides and magnet metal has migrated significantly towards East Asia with the relocation of manufacturing industries.

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Magnet and non-magnet REO demand, kt REO
Figure 198 Magnet and non-magnet demand (kt REO, source CRU)

Magnet REO demand composition, %

19.4. Supply, Geopolitics and Yttrium Export Controls

19.4.1. Chinese Dominance and Export Controls

China continues to dominate rare earth mining and processing, and it holds an overwhelming share of global yttrium separation capacity. USGS data and industry sources indicate that nearly all imported yttrium compounds and metal are derived from concentrates processed in China, with the United States relying on imports for 100% of its yttrium needs; approximately 93% of U.S. yttrium imports in recent years have come directly from China, with the remainder originating from material first processed there.

In April 2025, China imposed export restrictions on yttrium and six other rare earths in response to U.S. tariffs. These measures require exporters to obtain licences from Beijing and have:

Severely restricted shipments, with licences issued only for small parcels, and have caused long delays in delivery;
Stopped exports of yttrium to the U.S. entirely as of mid-2025, with overall exports to the rest of the world down about 30%.

Despite a subsequent partial relaxation of some rare earth curbs following high-level U.S.–China negotiations, the core April controls on yttrium remain in place, leaving access uncertain in the absence of a comprehensive political settlement.

19.4.2. Price Impacts and "Scramble for Yttrium"

The export controls have triggered what traders and analysts describe as a "scramble for yttrium":

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European prices for yttrium oxide have risen some 4,400% since January 2025 to ~USD270/kg.
Chinese domestic prices, by contrast, are quoted around USD7/kg, only modestly higher than at the start of the year and now trending down, reflecting controlled domestic supply and export bottlenecks rather than a lack of material in China itself.

Outside China, visible yttrium stocks have fallen sharply; some traders report inventories dropping from ~200 tonnes to ~5 tonnes, while others have run out entirely. Estimates of stocks outside China range from 1 to 12 months of consumption, varying significantly by company and sector.

Aerospace, energy and semiconductor producers are currently managing to avoid shutdowns, but industry participants warn that yttrium shortages are becoming a genuine chokepoint that increases production times and costs and could ultimately constrain output if sustained.

19.5. Recent Market Pricing and Peer Price Decks

19.5.1. Emerging Ex-China Price Divergence

Over the last several years, and especially since 2023–2025, the rare earth market has begun to exhibit a clear bifurcation between China-domestic prices and ex-China prices, particularly for Dy, Tb, Y, Nd and Pr. This divergence is driven less by geology or production cost than by policy, trade measures, ESG requirements and supply-chain strategy.

19.5.2. Drivers of Price Divergence

Key drivers include:

Export controls and licensing regimes
China's 2025 export restrictions on yttrium and six other rare earths have constrained the flow of material to the United States and other markets, even though domestic Chinese supply remains available. Licences have been granted only for small shipments, often with long delays, forcing ex-China buyers to bid aggressively for limited volumes.
Similar dynamics have applied at times to other magnet REEs where tighter export licensing, quotas or informal controls have altered trade flows and created two-tier markets.
Security-of-supply premiums
OEMs and governments in the U.S., EU, Japan and allied jurisdictions are increasingly willing to pay a premium for material sourced outside China or from Chinese supply chains that can demonstrate robust ESG and traceability standards.
This is reflected in long-term price decks used for ex-China projects such as Aclara's Carina and Northern Minerals' Browns Range, which assume Dy/Tb prices well above Chinese spot levels and explicitly contemplate ex-China premia in "divergence" scenarios.
Logistics, sanctions and trade-route risk
Sanctions regimes, tariffs, export controls and geopolitical tensions have increased freight, insurance and financing costs for certain trade routes.

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Buyers concerned about sudden policy-driven disruptions to Chinese exports are willing to pay more for material from stable, aligned jurisdictions, reinforcing price spreads.
ESG and certification requirements
Major OEMs (especially in automotive and defence) are subject to strict ESG, traceability and human-rights requirements and are therefore reluctant to rely solely on opaque supply chains.
Certified ex-China material can command a higher effective realized price once ESG and compliance costs are factored in.

19.5.3. Divergence Price Trends

Supply chain strategy and export licensing has created a “China price” and a much higher “rest-of-world price” for the same material.

Dy/Tb
Chinese spot prices for Dy and Tb (e.g., Dy oxide ~USD240/kg, Tb oxide~USD1,000/kg) represent domestic or FOB China values.
According to Benchmark Minerals, European markets for heavy rare earths have been facing significant pressure since April 2025 with prices for ex-China supply reaching $800USD/kg for Dy oxide and $3,625USD/kg for Tb oxide
Long-term assumptions used in PFS/DFS studies (e.g., Carina’s USD829/kg Dy oxide and USD3,056/kg Tb oxide) effectively embed an ex-China premium driven by supply-chain security.
Browns Range’s “divergence” scenario explicitly models a world in which ex-China prices diverge upward from China-domestic levels, producing a step-change in project NPV and IRR.
Nd/Pr
NdPr remain global commodities with relatively tighter arbitrage between China and ex-China, but the same forces — tariffs, export controls, strategic stockpiling and ESG filters — are pushing contract prices for ex-China Nd/Pr feedstock above China spot prices for long-term secure supply. This has been reinforced with the announcement of the investment in MP Materials by the US Department of War and their establishing a floor price of $110 USD/kg for NdPr oxides, almost double the China market pricing at the time.

The net result is that for Nd, Pr, Dy, Tb and Y, there is now:

A China price, reflecting domestic conditions and policy; and
An ex-China strategic price, reflecting scarcity of alternative sources, policy risk, ESG constraints and the willingness of OEMs/governments to pay for security of supply.

19.5.4. Implications for Current and Forecast Pricing for Lofdal

For Lofdal, the emerging bifurcation has several implications:

Price deck design
Base-case price assumptions should remain conservative, anchored to global consensus views and peer decks, but they should explicitly recognise that realised prices for Lofdal’s ex-China material may track ex-China strategic pricing.

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Upside cases model more pronounced ex-China premia for Dy, Tb, and Y and, to a lesser degree, Nd/Pr, similar to Browns Range's divergence scenario.
Basket value and by-product economics
Yttrium, historically treated as a relatively low-value bulk HREE, can no longer be assumed to trade at a small multiple of Chinese spot under all conditions. The 2025 experience shows that ex-China Y prices can spike to multiples of Chinese prices when export controls bite.
Lofdal's specifically heavy rare earths rich basket should therefore be evaluated not only on Dy/Tb value, but also on the option value of Yttrium as a strategic product in aerospace, semiconductors and gas turbines.
Offtake strategy and realised pricing
Long-term offtake agreements with ex-China separation and downstream partners should explicitly reference ex-China benchmark prices or indices, not purely China-domestic spot references.
For strategic customers (defence, aerospace, semiconductors, turbines), contract structures may include premiums over China-linked reference prices to reflect the value of traceable, non-Chinese Dy, Tb and Y supply.
Risk and sensitivity analysis
Future PFS/DFS work should include sensitivity cases where:
(a) China and ex-China prices converge (e.g., relaxation of controls), and
(b) divergence widens further (e.g., escalation of trade restrictions), with clear impacts on project NPV/IRR and Lofdal's competitive position.

From a marketing perspective, Lofdal is positioned as a provider of strategic ex-China Dy, Tb and Y supply into global mine-to-magnet and high-tech supply chains that operate on a divergence price curve.

19.5.5. Indicative Long-term View on Yttrium Prices

Lofdal Mixed Rare Earths Oxide product contains about 50% Yttrium oxide. No major peer PFS/DFS has yet published an explicit long-term yttrium price deck comparable to Dy/Tb assumptions. However:

The 4,400% spike to USD270/kg in Europe is assumed to be a crisis price rather than a sustainable long-term equilibrium.
Pre-crisis yttrium oxide prices were typically in the single digit to low double-digit USD/kg range.
The emergence of non-Chinese producers like ReElement suggests that, over time, some convergence towards intermediate price levels (well below crisis prices but above historical lows) is likely as new capacity comes online and policy support increases.

For Lofdal's planning purposes, yttrium should therefore be treated as a material value contributor with high strategic optionality: Base-case economics may use conservative Y price assumptions, with upside cases reflecting sustained premiums linked to ongoing geopolitical risk.

19.6. Marketing Strategy

Lofdal's marketing strategy explicitly recognises three main value pillars:

Heavy rare earths dopants: Dy/Tb
Classic magnet rare earths: Nd/Pr

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Speciality metal: Yttrium (Y) Key elements:
Product strategy
Produce a high-purity mixed REE carbonate with significant >60% in Dy, Tb and Y.
Collaborate with separation partners capable of producing separated Nd/Pr oxides, heavy REE oxides and high-purity yttrium oxide.
Target customers
Non-Chinese separation plants (e.g., Eneabba and other emerging hubs) and refining projects in North America, Europe and Japan.
Magnet manufacturers and OEMs in automotive, wind, industrial and defence sectors.
Aerospace, semiconductor and gas turbine OEMs concerned about yttrium supply security.
Government agencies and DFIs looking to de-risk critical mineral supply chains for defence and advanced manufacturing.
Contracting approach
Seek long-term offtake agreements that cover the full critical basket.
Structure contracts with price-floor, cost-plus or indexed pricing mechanisms that reflect the strategic nature of Dy/Tb and Y supply.
Use the JOGMEC JV to cooperate with Japanese industrial networks and allied governments on offtake-backed financing.

19.7. Existing Contracts

As of November 2025:

The JOGMEC Joint Venture remains the primary commercial framework, providing funding, strategic oversight and potential offtake rights for future REE production at Lofdal.
Standard service agreements are in place with drillers, laboratories and engineering firms.
No binding offtake or sales contracts have yet been executed for Nd/Pr, Dy/Tb or Y. However, in light of recent yttrium market developments, the Company anticipates prioritising early outreach to aerospace, semiconductor and turbine OEMs and their supply-chain partners.

19.8. Indicative Price Framework and Basket Value

For illustrative purposes only, the following indicative long-term price ranges (real 2025 USD) are proposed:

Rare Earth Oxide	Illustrative Long-Term Price (USD/kg)	Comment
La_{2}O_{3}		No value credited (by-product)
Ce_{2}O_{3}		No value credited (by-product)
Pr_{2}O_{3}	~100–167	In line with peer Nd/Pr price decks

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Rare Earth Oxide	Illustrative Long-Term Price (USD/kg)	Comment
Nd₂O₃	~103–160	In line with peer Nd/Pr price decks
Sm₂O₃	3–5	Minor value contribution
Gd₂O₃	78–131	Specialty applications
Tb₂O₃	2,450–4,030	Below some peer long-term cases, above current spot
Dy₂O₃	564–928	Range around Aclara and Northern Minerals assumptions
Y₂O₃	50–142	Reflects strategic premium vs historic levels, but below crisis prices of USD270/kg EU spot
Other HREEs Ho, Er, Tm, Yb, Lu	Various	Small, niche markets; priced case-by-case

19.9. Conclusions

Key conclusions from this updated market and contracts review are:

Demand for magnet rare earths (Nd/Pr/Dy/Tb) remains robust, and permanent magnets continue to account for the majority of rare earth value. Structural growth in EVs, e-mobility and wind are intact.
Yttrium has emerged as a critical bottleneck, with Chinese export controls in 2025 triggering a scramble for supply, a 4,400% price spike in Europe and heightened concern across aerospace, semiconductor and energy industries.
The global supply of Dy/Tb and Y remains heavily concentrated in China and Myanmar, with significant geopolitical and policy risk. Non-Chinese initiatives (Browns Range, Aclara, ReElement, Wicheeda) are important but insufficient on their own to fully de-risk supply.
Lofdal is well positioned as one of very few advanced major Dy/Tb- and Y projects outside China, with a supportive jurisdiction and a strategic JV partner (JOGMEC).
The Project's economic case is now underpinned by three value pillars: Nd/Pr magnet demand, Dy/Tb high-temperature magnet demand, and Yttrium demand from aerospace, semiconductors and turbine coatings. This diversified critical-material exposure increases the strategic attractiveness of Lofdal for OEMs and governments.
The rare earth market is characterised by a bifurcation between China-domestic and ex-China prices for Nd, Pr, Dy, Tb and Y. Export controls and security-of-supply concerns are creating a distinct ex-China strategic price regime for non-Chinese material. Lofdal's future products are expected to participate in this ex-China price environment.
Key residual risks are:
The durability and magnitude of ex-China premia for Dy/Tb and Y,
Future changes in Chinese export policy,
The speed and scale of new ex-China projects (including recycling/refining initiatives) entering the market and
Continued technological efforts to thrift Dy/Tb and substitute or optimise Y usage.

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These will be addressed through scenario analysis in the PFS/DFS and by prioritising offtake structures and financing solutions that recognise the strategic nature of Lofdal's product suite.

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20. Environmental Studies, Permitting and Social or Community Impact

William van Breugel P.Eng. (SGS) has relied upon SLR Consulting (Namibia) (Pty) Ltd (SLR), who completed an independent analysis of permitting and environmental requirements for Item 20.

The Environmental Studies, Permitting and Social or Community Impact section presented below was completed in April 2024, before the value engineering for the PFS.

20.1. Project Background

Namibia Rare Earths (Pty) Ltd (NRE), a subsidiary of Namibia Critical Metals Incorporated (NMI), is a Canadian company listed on the Toronto Stock Exchange (TSXV: NMI OTC: NMREF) Venture Exchange with a diverse project portfolio in Namibia. NRE holds Mining Licence (ML) 200 on Farm Lofdal, which falls within the //Huab and Doro!Nawas Conservancy Areas in the Kunene Region of Namibia (please refer to Figure ES-1-1 and Figure ES-1-2. In 2016, SLR Environmental Consulting [Namibia] (Pty) Ltd (SLR) undertook three Environmental Impact Assessments (EIAs) in relation to the Project1, namely:

The EIA for the construction and operation of the powerline for the Project – approved by Namibia Power Corporation (NamPower) (SLR (a), 2016);
The EIA for the water pipeline for the Project – approved by NamWater (SLR (b), 2016); and
The EIA for the Lofdal Mining Project (SLR (c), 2016).

Following the EIA process and review ((SLR (c), 2016), NRE was granted an Environmental Clearance Certificate (ECC) in December 2017 by the Ministry of Environment, Forestry and Tourism (MEFT) to support the Mining Licence (ML) application. The ECC was renewed in May 2021 and is valid for a period of three (3) years. A corresponding Mining Licence 200 (ML200) was issued by the Ministry of Mines and Energy (MME) in May 2021.

Under the provisions of the Environmental Contract, NMI is required to submit bi-annual environmental reports detailing work and potential impacts. NMI has fully complied with this provision and copies of these reports for ML 200 are filed in company files which are complete and up to date. The Environmental Clearance Certificate (ECC) for ML-200 was originally issued December 05, 2017 (for the Mining License application) and was subsequently renewed in 2021 and 2024. The latest renewal is dated April 24, 2025, and is valid until April 24, 2028.

Copies of the existing ECCs are included in Appendix A.

Since the granting of the ECC and the Mining Licence for the Lofdal Mining Project, subsequent test work undertaken by NRE has indicated that the resource of the mine is significantly larger than originally outlined in the 2016 EIA (SLR (c), 2016). It follows that the ECC for the Lofdal Mining Project needs to be amended to cater to a larger mining operation and additional beneficiation steps (hereafter referred to as the Project).

It must be noted that only exploration related work and pilot scale mining are currently taking place on site and as such, no infrastructure relating to the Lofdal Mine has been established.

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1 Collectively refers to the three existing approved EIA’s that make up the Lofdal Mining Project.
2 Pilot scale mining refers to a phase in the development of a mining project where a small-scale operation is established to test and demonstrate the feasibility of full-scale production.

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Figure 199 Regional Setting

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Figure 200 Local Setting

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20.2. Motivation for the Project

The mining sector is a cornerstone of Namibia's economy, contributing significantly to GDP, employment, and government revenue. The Project, led by NRE, aims to develop a high-grade rare earth element (REE) mine and processing plant, capitalizing on the global demand for critical minerals such as dysprosium and terbium.

The Project, with an estimated capital investment of USD 207 million, is expected to generate USD 1.1 billion in pre-tax revenue over its 16-year lifespan, contributing USD 400 million to Namibia's economy. It will create approximately 300 jobs during construction and 243 permanent positions during operations. Additionally, it includes a 5% shareholding for historically disadvantaged Namibians, local procurement initiatives, and continued support for community programs, including education, water access, and social welfare.

While the Project will have environmental impacts, a revised Environmental Management Plan (EMP) is being developed to align with international best practices. The EMP will ensure sustainable mining operations by mitigating negative effects and implementing adaptive management strategies.

This project represents both a strategic economic opportunity and a step toward diversifying the global REE supply chain while supporting Namibia's Vision 2030.

20.3. Project Description Overview

NRE plans to develop an open-pit mine and processing plant at Farm Lofdal to produce rare earth element (REE) carbonate. The Project will use conventional open-cast mining methods across two pits (Area 2B and Area 4). After site preparation, drilling and blasting will loosen waste rock and ore, which will be transported using truck and shovel methods. Waste rock will be stockpiled in designated dumps, while topsoil will be preserved for ongoing rehabilitation.

Onsite processing will include a concentrator, refinery, and various treatment stages. The concentrator will handle primary crushing, milling, and flotation, with recovered water being recycled. The refinery will extract, purify, and precipitate REEs through sulphuric acid bake, leaching, impurity removal, and drying. The final REE product (99% grade TREO carbonate) will be packaged for export. Additionally, a minor amount of uranium present in the ore will be extracted through ion exchange and precipitated separately.

The Environmental Clearance Certificate (ECC) initially approved in 2016 covered the mine's development. However, due to an expanded mining operation and additional processing steps, an amendment application was submitted to MEFT on October 19, 2023. Following this, specialist studies on various environmental and socio-economic aspects were conducted, leading to revisions in the mine layout plan. These updates ensure alignment with engineering assessments and environmental recommendations. A comparative overview of the approved 2016 EIA and the updated 2023 layout is provided in Table 125.

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Table 125 Project Data Summary that provides perspective on the scale of and amendments made to the Project

Group	Specific	2016 Project Details	Project Details (2023)	Amended Project Details (2024)
Mining	Target mineral	Rare Earths	Rare Earths	Rare Earths
	Mineable area	Main pit: 40 hectares
Smaller satellite pits up to 10 km away within the ML area may be developed in the future	There will be 2 open pits (A4 and A2B) with a total footprint of 68.38 ha (A4 open pit: 48.87 ha and A2B open Pit: 19.51 ha)	There will be 2 open pits (A4 and A2B) with a total footprint of 78.81 ha (A4 open pit: 65.36 ha and A2B open Pit: 13.45 ha)
	Depth of the minerals below surface	The minerals occur at the surface and will be mined to a depth of 200 m	The minerals occur at surface and will be mined to a depth of 300 m	The minerals occur at the surface and will be mined to a depth of 140 m at Area 2B pit and 330 m at Area 4 pit
	Rate	840 000 tonnes/annum of ore to be sent to the crusher	2 000 000 tonnes/annum of ore to be sent to the crusher	2 160 000 tonnes/annum of ore to be sent to the crusher
	Life of mine	7 years	16 years	16 years
	Extent of areas required for infrastructure	The processing plant, pit, tailings storage facility (TSF), waste rock dump (WRD) and solar plant would cover about 15 km² (1 500 ha).	The processing plant, pits, tailings storage facility (TSF), WRDs and solar plant would cover about 20 km² (2 000 ha).	The processing plant, pits, tailings storage facility (TSF), WRDs and solar plant would cover about 21 km² (2 100 ha).
Mine residues	Waste rock: the mine material that does not contain rare earths to be processed will be stockpiled	Two alternatives of one large WRD, or two smaller WRDs located to the south of the A4 pit. The WRD (s) would have an end of mine storage capacity of 50 000 000 m³.	Two WRDs located northwest of the A2B open pit and south of the A4 open pit. The South WRD and A2B WRD WRDs will have an end of mine storage capacity of 43 797 500 m³ and 11 718 933 m³, respectively.	Three WRDs located northwest of the A2B open pit and north and south of the A4 open pit. The A4 WRDs (north and south) and A2B WRD will have an end of mine storage capacity of 65 835 203 m³, 34 321 380 m³ and, 10 834 798 m³ respectively.
Processing	Process Plant	Mill Process Plant	Mill Process Plant, with a hydrometallurgical plant added to the processing plant	Mill Process Plant, with a hydrometallurgical plant added to the processing plant
	Flotation Plant	Flotation Plant with throughput of 0.9 Mt/a	Flotation Plant with throughput of 2.1 Mt/a	Flotation Plant with throughput of 2.16 Mt/a

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Group	Specific	2016 Project Details	Project Details (2023)	Amended Project Details (2024)
Product	Final product for export	low-grade mineral concentrates (xenotime concentrate)	99% grade TREO product	99% grade TREO product
Processing residues	Tailings	The TSF had a capacity to store 3.24 million tonnes (Mt), with a maximum footprint of 5.3 ha.	The TSF has a capacity to store 26.8 Mt, with a maximum footprint of 137 ha.	The TSF has a capacity to store 26.8 Mt, with a maximum footprint of 227 ha.
Resource use	Water demand	Approximately 1 000 000 m³ per year	Approximately 1 000 000 m³ per year	Approximately 1 600 000 m³ per year
	Power demand	21.178 million kwh/annum	40.951 million kwh/annum	85.989 million kWh/annum
Employment	Staff: Construction	Approximately 300	Approximately 300	Approximately 300
	Staff: operational	Approximately 226	Approximately 243	Approximately 243
	Operating times	24 hours a day, 7 days a week	24 hours a day, 7 days a week	24 hours a day, 7 days a week

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Figure 201 Site layout (Approved 2016 layout and amended (current) layout

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20.4. Project Alternatives

The key alternatives considered in the EIA process include:

Project Layout Alternatives: The initial 2023 site layout was revised in 2024 following specialist assessments. Key changes include relocating the Area 2B Waste Rock Dump (WRD) and topsoil stockpile to ensure Road D2625 remains unobstructed. However, the processing plant remained on a ridgeline for operational efficiency. The updated layout also reduced impacts on biodiversity and water drainage.
Return Water Dam (RWD) Pipeline Alternatives: Two routing options were considered for the pipeline connecting the processing plant to the RWD. Option 1, a 2.9 km route, crosses a haul road, while Option 2, a 3.56 km route, follows an access road before turning northward.
Alternative Land Uses: Post-mining rehabilitation options include restoring the land for livestock farming, tourism, or other mining-related activities. Eco-tourism potential is being explored, with final decisions to be made in consultation with stakeholders.
No-Go Option: If the Project does not proceed, Namibia, and the local communities will forgo economic benefits such as job creation, infrastructure development, and increased revenue. While this option would avoid environmental impacts, the additional 16-year mineral exploitation period would generate significant economic growth and outweigh potential negative impacts, provided proper mitigation measures are implemented.

20.5. EIA and Public Consultation Process

The EIA process for the Project is regulated under Namibia's Environmental Management Act (EMA), 2007, and its associated regulations. The process ensures that potential environmental and social impacts are identified and mitigated to achieve sustainable project development.

The key phases of the EIA process include:

Project Initiation Phase: A desktop review was conducted to identify key environmental risks and necessary specialist studies. The Project was registered with the MME and the MEFT. MEFT issued a screening notice requesting an Environmental Management Plan (EMP).
Scoping Phase: This phase involved stakeholder engagement, baseline environmental assessments, identification of key impacts, and determination of project alternatives. The scope of the EIA, including methodology and specialist study requirements, was also defined.
Impact Assessment Phase: A comprehensive evaluation of potential environmental and social impacts, including biophysical and socio-economic factors, was conducted. The assessment incorporated specialist studies and an impact rating methodology to determine significance, reversibility, and mitigation measures.
The findings of the EIA, along with the EMP, will inform the decision-making process for the ECC amendment application. The Project's impacts and mitigation measures are designed to ensure environmentally responsible and sustainable mining operations.
The PCP is being undertaken per the requirements of Regulation/Part 21 of the EMA as follows:

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Project Initiation Phase: During this phase, Interested and Affected Parties (I&APs) were notified of the Project and invited to provide comments. No issues were raised. Stakeholder communications are recorded in Appendix F.
Scoping Phase: Consultations with authorities were ongoing. The scoping phase commenced after MEFT issued a screening notice requiring a Scoping Report and an Environmental Management Plan (EMP) for the ECC Amendment Application. I&APs had the opportunity to review the Draft Scoping Report from 7 December 2023 to 10 January 2024.
EIA Phase: The Draft EIA Report was open for public comment from 10 March 2025 to 8 April 2025. The registered I&APs received notifications via email, along with a revised Background Information Document (BID) and Non-Technical Summary (NTS). Once the review and comment period is completed, stakeholder feedback is compiled into an updated Comments and Responses Report (CRR) for inclusion in the Final EIA Report.
The Final EIA Report is submitted to MEFT and MME for review and decision- making.
Authority Decision Phase: The MEFT will issue a decision on the ECC Amendment Application based on the Final EIA Report and EMP. SLR will notify all registered I&APs of the decision and the appeals process. Further consultation with MEFT and MME may take place as needed until a final decision is reached.

20.6. Profile of the Receiving Environment

20.6.1. Biophysical Environment

The baseline biophysical environment of the Project area encompasses key environmental components and is summarised in Table 126.

Table 126 Summary of the Profile of the Biophysical Baseline Environment

Aspect

Description

Climate and Meteorology.

The Project area, located in north-western Namibia, experiences an arid climate with annual temperatures averaging above 22°C, ranging from a maximum of 40.0°C in October to a minimum of -0.4°C in June. Rainfall is scarce, with an annual mean precipitation of 229 mm, primarily occurring from January to March, while evaporation significantly exceeds precipitation at 2,380 mm per year, leading to a substantial environmental water deficit. Wind patterns predominantly come from the south-west and north-east, influencing dust dispersion and air quality.

Topography, Soils, and Land Capability

The Project is situated on moderately elevated, undulating terrain, with prominent ridgelines to the north and west. Mining infrastructure may alter natural drainage patterns, potentially impacting water flow and landscape aesthetics. The primary soil type, Eutric Skeletic Lithic Leptosol, is shallow (<10 cm depth), highly stony, and has limited agricultural potential due to poor water retention and susceptibility to erosion. The land capability is restricted, supporting only low-density livestock and wildlife

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Aspect	Description
Surface Water and Hydrology	The Project area lies within the Huab River catchment, with ephemeral tributaries like the Sout River shaping local drainage patterns. Mining operations, particularly waste management infrastructure, may impact surface water quality through runoff and seepage. Hydraulic assessments indicate that only a few river sections maintain a groundwater contribution, emphasizing the need for stringent water management strategies.
Groundwater and Geology	The Lofdal project is located within a hydrogeological framework characterized by low-yield fractured aquifers with limited recharge potential. Groundwater levels range between 5 and 50 meters below the surface, with flow directed towards the Atlantic Ocean. The region’s groundwater is classified as brackish, with elevated total dissolved solids (TDS) and localized uranium and metal exceedances. The geology comprises metamorphic basement rocks, including carbonatite intrusions, which are associated with rare earth element (REE) mineralization.
Biodiversity and Ecology	The Project falls within the Acacia Savanna Biome, dominated by Mopane and Thornbush vegetation, with moderate plant species richness (291 species). Five species of conservation concern were recorded, including protected trees like Combretum imberbe and Commiphora multijuga. The area supports diverse vertebrate fauna, including 63 mammal species, 138 bird species (including Red Data- listed raptors), and 46 reptile species. Key habitats include rocky hills, large watercourses, and drainage lines, with several zones classified as highly sensitive due to their ecological importance.
Air Quality and Noise	Baseline air quality data indicate low dustfall rates except in September when thresholds were exceeded due to wind activity. PM10 and PM2.5 levels showed frequent exceedances of health-based guidelines, largely due to natural dust and minor on-site activities. Noise levels in the Project area remain within acceptable limits, with primary sources including wind, local wildlife, and occasional human activities.
Radiation and Public Health Considerations	The presence of naturally occurring radioactive materials (NORM), including uranium and thorium, is characteristic of the REE mineralization in the Project area. Baseline radiation assessments indicate background levels comparable to global averages, with total exposure estimated at 2.0 mSv/a. Future mining operations will require active radiation management to mitigate risks to workers and nearby communities.

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20.6.2. Cultural Heritage

The Project area has a cultural and heritage landscape, encompassing archaeological, historical, and palaeontological resources that are summarised in Table 127.

Table 127 Summary of the Profile of the Cultural Heritage Baseline Environment

Aspect	Description
Archaeological and Cultural Heritage Resources	The Project area features deeply weathered metasediments, isolated carbonatite outcrops, and volcanic formations, with prominent sandstone mesas to the west. Archaeological evidence indicates human occupation spanning at least 5,000 years, including small rock shelters with rock art, stone-based hut encampments, grave cairns, and surface artefact scatters. Twyfelfontein, a UNESCO World Heritage Site, is located nearby and is renowned for its rock engravings. Additionally, historical remains from colonial-era farming settlements, including graves and abandoned structures, are present.
Four key cultural heritage sites have been identified within the Project area:
• Old Burial Site – A colonial-era farm graveyard with four graves, including a marked grave from 1955.
• Abandoned Residential Structure – Likely dating to the colonial era, now in a state of disrepair.
• Animal Water Structures – An old water tank and drinking trough.
• Livestock Drinking and Dipping Area – A colonial-era structure used for water storage and animal disease treatment.
Palaeontological Resources	The area is also of some palaeontological interest, most finds being associated with the Jurassic sediments including local concentrations of petrified wood (the National Monument Petrified Forest site, proclamation 4/50, is situated close to the southern margin of the lease area).

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20.6.3. Socio-Economic

The socio-economic baseline of the Project provides an overview of the region’s demographic, economic, infrastructure, and social conditions and is summarised in Table 128.

Table 128 Summary of the Profile of the Socio-economic Baseline Character

Aspect	Description
Visual Baseline	The visual character of the Project area is shaped by its landscape, scenic quality, sense of place, and key visual receptors. The Project area is defined by prominent landscape features, including the town of Khorixas, located 25 km southeast, and sparsely populated farmlands used for livestock rearing. The closest major road is the C39, which is significant for tourism due to its proximity to the Petrified Forest, a national monument containing the ancient Welwitschia mirabilis plant. Other key tourism sites in the region, such as the Aba-Huab Lodge and Twyfelfontein rock engravings, are located outside the mine’s immediate visual impact zone. Small rural hamlets related to the //Huab and Doro !Nawas Conservancies also exist, requiring careful management to minimize visual intrusion.
The region’s low vegetation height offers minimal screening, making the landscape highly susceptible to visual alterations from mining activities. As the area has not previously been an industrial or mining landscape, the introduction of large-scale mining infrastructure will be highly visible, particularly from elevated areas. The topography, which includes rocky outcrops, adds significant scenic value, especially for eco-tourism, and should be considered for conservation partnerships with local conservancies.
Traffic	The existing road network around the Project consists of a combination of main, district, and local roads that facilitate transportation for mining operations, local communities, and economic activities. The primary access road is the M0126, which connects Khorixas and Uis and is essential for regional connectivity. District roads D2633 and D2625 provide direct access to the mine, while an extensive network of gravel roads supports transportation within the mining site.
Land Uses	The primary land uses around the Project are subsistence livestock farming and wildlife-based economic activities, including trophy hunting and tourism. However, the area has a history of mining activities, with previous environmental damage recorded due to dimension stone quarrying. Seven areas have been identified where open-pit quarrying, stockpiling, and road construction caused land disturbance, with no rehabilitation efforts undertaken. Additionally, several tantalite mining claims exist in the northern part of the Project area, but mining activities ceased in 2014 due to declining resource quality. Some rehabilitation efforts have been attempted, but many hand-made diggings remain open. Another environmental impact assessment has been conducted for further small-scale mining in the region.

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Aspect

Description

Socio-Economic Structure and Profile

The Project falls under the Kunene Regional Council, which aligns its development plans with Namibia’s Vision 2030 strategy, focusing on poverty alleviation, job creation, healthcare, and sustainable resource management. The regional governance structure supports decentralization, with various ministries operating under the regional council. However, Outjo is the only municipality in the region capable of independently financing its development projects, with other areas reliant on government grants.

The Khorixas Constituency, where the Project is located, has a population of approximately 12,500 people spread across 22,000 square kilometres, resulting in a low population density of 1.76 people per square kilometre. The town of Khorixas accounts for more than half of the population, with smaller settlements scattered throughout the constituency. The average household size is four people, and informal settlements have increased due to rural-urban migration. Traditional authorities play a significant role in land administration, with the #Aodaman Traditional Authority overseeing the Lofdal area, while the Swartbooi Traditional Authority is located nearby. Many of the Project’s workers are expected to originate from these traditional groups. Employment opportunities in the region are limited, with a high unemployment rate of 50% and youth unemployment reaching 53% (65% for those under 24). Agriculture remains the largest employment sector, engaging 53% of the labor force, primarily in subsistence farming. The mining sector is currently small, employing only 2% of workers. The region also has a well-developed tourism sector, with nearly half of Namibia’s conservancies located in Kunene. Wildlife-based tourism is a key economic driver, and conservancies such as //Huab and Doro !Nawas have established joint ventures with tourism operators. However, the //Huab Conservancy lost its primary income source when a mining license was granted in its exclusive wildlife area, causing the closure of its award-winning lodge.

Infrastructure in the region is underdeveloped, with 67% of rural households having access to safe water, but 63% lacking proper sanitation. Electricity access is limited to 31% of households, with the majority relying on wood or charcoal for cooking. Education levels are also low, with 40% of residents leaving school after completing primary education, and only 14% attaining secondary education. The literacy rate in Khorixas is 84%, below the national average of 89%. Health services are limited, with the nearest facilities located in Khorixas and Fransfontein. The region has Namibia’s highest teenage pregnancy rate (39%) and low contraceptive use (52.2%), which could be exacerbated by an influx of mine workers. Additionally, the Kunene Region reports high levels of gender-based violence, with 36% of women experiencing physical violence since the age of 15.

The closest sensitive receptors to the mining project are the Oas Post 3 homestead and the Lofdal Homestead, both of which are expected to be relocated due to potential air quality, noise, and radiological impacts. |

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20.7. Specialist Studies

The assessment has been informed by extensive specialist studies on key environmental aspects previously considered in the 2016 EIA, including groundwater impact assessment, heritage and archaeological assessment, visual impact assessment, biodiversity impact assessment (covering avifauna, vertebrates, flora, and invertebrates), social impact assessment, noise impact assessment, radiation assessment, and air quality impact assessment. These studies have been updated to reflect the revised layout plan.

20.8. Quantification of Impacts

The anticipated impacts associated with the Project were assessed according to SLR's standardised impact assessment methodology. The impact assessment methodology enables the assessment of biophysical, cultural, and socio-economic impacts including cumulative impacts and impact significance through the consideration of intensity, extent, duration, and the probability of the impact occurring.

The assessment indicates that the Project's impacts range from highly positive to highly negative. Notably, the assessment assumes that the group of residents currently residing near the Project (specifically at Oas Post 3 and the Lofdal Homestead, within the vicinity of the Area 2B WRD and TSF) will be relocated prior to the commencement of mining construction activities. This relocation is essential, as their current living conditions would become uninhabitable once these activities commence. The assessment has determined that the potential health and safety risks to these residents are significant.

A summary of the EIA is provided in Table 129.

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Table 129 ES 1-5: Summary of the Findings of the EIA

	Issue	Relevant Project Phase	Significance Rating Impact Assessment
					2016 EIA Rating		2024 EIA Rating
		Construction	Operation	Decommissioning and Closure	Unmitigated Scenario	Mitigated Scenario	Unmitigated Scenario	Mitigated Scenario
BIOPHYSICAL	Loss of Soil and Land Capability Through Contamination	X	X	X	Medium	Low	Medium	Low
	Loss of Soil and Land Capability Through Physical Disturbance	X	X	X	High	Medium	High	Medium
	Contamination of Surface Water Resources that could influence availability for third-party supply		X	X	High	Low	Medium	Low
	Alteration of Natural Drainage Patterns Affecting Natural Run-off to the Catchment and Third-Party Use	X	X	X	Medium	Low	Medium	Low
	Alteration of Natural Drainage Patterns Attributing to Flooding on Site	X	X	X	Impact Not Identified	Impact Not Identified	Medium	Low
	Contamination of groundwater that could influence availability for third-party supply			X	Medium	Low	Medium	Low
	Lowering of Groundwater Levels that can Influence	X	X		Medium	Low	High	Low

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Issue	Relevant Project Phase			Significance Rating Impact Assessment
	Construction	Operation	Decommissioning and Closure	2016 EIA Rating		2024 EIA Rating
				Unmitigated Scenario	Mitigated Scenario	Unmitigated Scenario	Mitigated Scenario
Availability for Third-Party Supply
Lowering of Groundwater Levels Affecting Baseflow				Impact Not Identified	Impact Not Identified	Negligible	Negligible
Pollution emanating from seepage of runoff from non-mineral wastes and processing plant which have a negative impact on groundwater quality				Impact Not Identified	Impact Not Identified	Negligible	Negligible
Physical Loss of Biodiversity	X	X	X	High	Low	High	Low
Interference with the crucial ecological process of water flows during Operation		X	X	High	Low	High	Low
General disturbance of biodiversity	X	X	X	High	Low	High	Low
Rehabilitation measures, and ecological restoration			X	Impact Not Identified	Impact Not Identified	Low	Low
Increase in disturbing noise levels affecting sensitive receptors	X	X	X	Low	Low	High	Medium
Atmospheric Emissions: increase in dustfall, PM2.5 and PM10 during construction	X			Medium	Low	Medium	Low

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Issue	Relevant Project Phase			Significance Rating Impact Assessment
	Construction	Operation	Decommissioning and Closure	2016 EIA Rating		2024 EIA Rating
				Unmitigated Scenario	Mitigated Scenario	Unmitigated Scenario	Mitigated Scenario
Atmospheric Emissions: increase in PM2.5 during operations		X		Medium	Low	Medium	Low
Atmospheric Emissions: increase in PM10 during operations		X		High	Medium	High	Medium
Atmospheric Emissions: increase in dustfall during operations		X		Medium	Low	Medium	Low
Atmospheric Emissions: increase in dustfall PM2.5 and PM10 during closure			X	Low	Low	Low	Low
Direct exposure to gamma radiation from on-site radioactive materials	X	X	X	Negligible	Negligible	Negligible	Negligible
Inhalation of radon and radon progeny	X	X	X	Negligible	Negligible	Negligible	Negligible
Direct ingestion of radioactively contaminated soil or dust	X	X	X	Negligible	Negligible	Negligible	Negligible
Direct ingestion of radioactively contaminated water	X	X	X	Negligible	Negligible	Negligible	Negligible

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	Issue	Relevant Project Phase			Significance Rating Impact Assessment
		Construction	Operation	Decommissioning and Closure	2016 EIA Rating		2024 EIA Rating
					Unmitigated Scenario	Mitigated Scenario	Unmitigated Scenario	Mitigated Scenario
	Indirect ingestion of radioactively contaminated soil or water	X	X	X	Negligible	Negligible	Negligible	Negligible
	Inhalation of radioactive atmospheric dust	X	X	X	Low	Low	Low	Low
	Blasting Impacts affecting sensitive receptors and third-party infrastructure	X	X		Medium	Low	Medium	Low
CULTURAL HERITAGE	Loss or Damage to Archaeological Resources (Sites 2)	X	X	X	Medium	Low	Medium	Very Low
	Direct Loss or Damage to Archaeological Resources (Sites 1)	X	X	X	Not assessed		Low	Insignificant
SOCIO-ECONOMIC	Landscape Degradation due to the Project	X	X		High	Medium	High	Medium
	Increased traffic	X	X	X	High	Medium	High	Medium
	Changes in Land Use Affecting Livelihood and Resulting in Loss of Land	X	X	X	Impact Not Identified	Impact Not Identified	High	Medium

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Issue	Relevant Project Phase			Significance Rating Impact Assessment
	Construction	Operation	Decommissioning and Closure	2016 EIA Rating		2024 EIA Rating
				Unmitigated Scenario	Mitigated Scenario	Unmitigated Scenario	Mitigated Scenario
Impact on Revenue Generation by Conservancies Due to Changes in Land Use	X	X		Impact Not Identified	Impact Not Identified	Medium	Low
Economic Benefit to National and Local Economies	X	X	X	High	High	Very High	Very
Job Creation and Skills Development	X	X	X	High	High	Very High	Very High
In-migration impacts on local communities and service provision	X	X	X	High	Medium	High	Medium
Hazardous Excavations and Infrastructure Resulting in Safety Risks to Third Parties and Animals	X	X	X	High	Low	High	Low
Mine decommissioning and closure			X	High	Medium	High	Medium

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20.9. Mitigation

20.10. Environmental Statement and Conclusion

It is anticipated that it will be possible to successfully mitigate all of the environmental impacts to acceptable levels and the implementation will be monitored and audited to determine the effectiveness of the measures implemented.

No fatal flaws/aspects have been identified that could render this the Project unfeasible and impractical. Therefore, it is SLR’s opinion that, based on the findings of the EIA process, there is no reason why the proposed development may not continue subject to the implementation of recommended mitigation measures. The Project should be allowed to proceed, considering the positive social and economic benefits associated with the Project.

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21. Capital and Operating Costs

21.1. Basis of Estimate

The accuracy of cost estimates for this PFS is ± 30%.

The cost base is an engineering, procurement and construction management (EPCM) model for process and non process facilities and contractor operated / owner managed operations for mining. Costs are estimated using a combination of vendor quotes, base commodity prices and established factors suitable for a Prefeasibility level assessment.

Table 130 provides the assumptions and basic commodity prices used for the basis of estimate that cover all functional areas.

Exchange rates are based on a 2-year trailing average with the record dates being September 2023 to September 2025.

Diesel Fuel prices are based on Namibian prices for July 2025, with reductions for tax refunds and consumption discounts.

Table 130 Base Criteria for Project CAPEX and OPEX

Base Parameters for Cost Estimation		Units
Exchange Rates	USD:ZAR	$	$0.055
	ZAR:USD	ZAR	ZAR 18.23
	USD:NAD	$	$0.055
	NAD:USD	NAD	NAD 18.23
Process Feed Parameters	Annual Tonnes to Crusher	tonnes	3 010 000
	High Grade Feed to Plant	tonnes	1 100 000
	Low Grade Feed to Ore Sorter	tonnes	1 9100 00
	Total Feed to Process Plant	tonnes	1 482 000
Base Commodities	Diesel Fuel	USD / l	$0.87
	Water Consumption Cost	USD/m³	$0.326
	Power Cost (Blended)	USD / kWh	$0.134
Transport and Smelting Costs	Concentrate Transport Costs	per tonne concentrate	$42.15
	Metal Separation LREO	USD per kg	$2.00
	Metal Separation HREO	USD per kg	$8.00

No provision has been made for cost escalation or exchange rate fluctuations.

21.2. Project Capital Summary

Overall Project capital is estimated at USD347.9M, inclusive of contingency.

Initial capital is estimated at USD273.4M. Contingency for initial capital is estimated at USD54.9M. Sustaining capital is estimated at USD17.2M with contingency of USD2.5M

Table 131 provides the capital breakdowns by key areas. Individual breakdowns are provided in respective sections below.

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Table 131 Capital Summary (USD)

Area	Total	Initial Capital	Sustaining Capital
Mining	$27 620 316	$21 927 081	$5 693 235
Process	$181 571 767	$181 571 767
Tailings	$21 590 251	$11 068 059	$10 522 192
Non Process Facilities	$58 746 582	$58 746 582
Closure	$1 039 987	$ 73 564	$ 966 423
Total Capital without Contingency	$290 568 903	$273 387 053	$17 181 850
Contingency	$57 360 067	$54 859 391	$2 500 676
Total Capital with Contingency	$347 928 971	$328 246 444	$19 682 526

21.2.1. Mining

Cost estimates for mining capital are based on the mine plan details provided in Chapter 16, and contractor quotes based on the mine plan. Contractor hourly rates for mobile equipment are quoted on a capital recovery basis, thus there is no initial capital for the mobile mining fleet, except for equipment mobilization.

The vendor quotes also include costs for the contractor's fixed plant equipment:

Mining Contractor Yard – Laydown Area
Mining Contractor Yard - Water Reticulation
Mining Contractor Yard - Power Reticulation
Offices / Admin Block
Change Houses / Rest Rooms
Mine Store / Warehouses
Light vehicle Workshop
HME Workshop & Equipment Overhaul Paddy
Welding Workshop / Bay
Wash Bay
Fuel & Lubrication Bay
Tyre Bay
Explosive Yard
Magazines
Water Truck Filling Point
Sewerage and Water Treatment Plant
Drilling OEM facility

Mining costs also include owner indirects for start up and initial pre-stripping. The capital costs for owner indirects include:

Owner's Mining Labour during pre-production stripping
Light Vehicles (4x4 D/C) capital and usage
Laptops & Screens
Office Equipment
Geology, Mine Planning and Survey Software
GPS, Total Stations and Drones
Truck Dispatch System Software / Network
Fuel Management System (by Fuel Supplier)

The mining capital cost breakdowns are provided in Table 132.

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Table 132 Mining Capital Cost Breakdown (USD)

Mining Capital	Total	Initial Capital	Sustaining Capital
Contractor Mobilisation and Demobilisation	$3 223 045	$1 387 321	$1 835 724
Contractor Fixed Plant Capital	$5 722 975	$5 722 975
Contractor Staff Mobilization	$ 314 050	$ 284 592	$ 29 459
Contractor Pre-Production Mining Costs Variable	$10 628 724	$10 628 724
Contractor Pre-Production Mining Costs Fixed	$2 014 828	$2 014 828
Owner's Pre-Production Mining Costs (Indirects)	$ 976 340	$ 976 340
Owner's Mining Capital	$4 740 354	$ 912 301	$3 828 053
Total Mining Capital	$27 620 316	$21 927 081	$5 693 235

21.2.2. Process Capital

21.2.2.1. Estimating Approach

The purpose of the capital cost (Capex) estimate is to provide the Prefeasibility Study capital cost for the Lofdal Rare Earths Project in Namibia.

The guidelines and procedures for preparing the process capital cost estimate is in accordance with the SGS Bateman Estimator's Best Practice Guide PCNG-0920-002 Rev 1 definition for a Prefeasibility estimate with an estimating accuracy of ±30% which acceptably meets the requirements of a Class 3 Prefeasibility study.

The capital cost estimate is structured in accordance with the project work breakdown structure (WBS) and a mechanical equipment list (MEL) quantified and specified in sufficient detail to identify all major equipment in accordance with the scope of works and preliminary design information.

21.2.2.2. Qualifications, Assumptions, Exclusions and Exceptions

SGS Bateman has solicited pricing for the major mechanical packages by expediting budget quotations and some informal email budget quotations where the vendors could not provide a formal quotation. The procurement team has approached multiple vendors and specialist turnkey suppliers for each mechanical equipment package.

The budget quotations received were technically analysed by the engineering team and recommendations were made based on the quality of information provided by the vendors in accordance with the various packages that have been issued. The recommended vendors were then commercially adjudicated by the procurement team for compiling budget costs for the equipment packages.

Costs for transportation of goods to the site and vendor assistance have been factorised for the packages which did not have the transport and vendor assistance included in the proposals received from the respective vendors. All other disciplines have been factorised from the mechanical equipment costs based on the factors and norms obtained from recent similar project completed by SGS Bateman.

21.2.2.3. Estimating Criteria

21.2.2.3.1. Capital Cost Estimate

The estimating approach undertaken is by identifying and quantifying the mechanical equipment from the MEL which included the platework and conveyors then pricing the supply component of these items using a combination of vendor budget quotations, and in-house database prices. The erection cost of these items factored as a percentage cost of the supply component.

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The cost of executing the bulk earthworks and civil works is priced as a percentage factored from the cost of the supply of mechanical equipment. The balance of the direct field cost is priced as a factored cost of the mechanical equipment supply with the preliminary and general costs of the bulk earthworks, civil works, structural mechanical platework & piping (SMPP) and electrical & instrumentation (E&I) being priced as additional costs using a percentage factor of their respective sections.

All the percentage factors used on the capital estimate have been based on database information compiled from previous experience and norms of recently completed similar projects by SGS Bateman.

21.2.2.4. Estimating Accuracy

The estimate was compiled in accordance with the SGS Bateman Estimator's Best Practice Guide PCNG-0920-002 Rev 1 definition for a Prefeasibility estimate with an estimating accuracy of $\pm 30\%$ which acceptably meets the requirements of a Class 3 Prefeasibility study.

21.2.2.5. Base Date

The base date of the capital cost estimate is September 2025. Forward escalation and exchange rate variations have been excluded.

21.2.2.6. Scope Definition

Process and discipline engineering documentation utilized in compiling the estimate include the following:

Mechanical Equipment List (MEL) for major equipment to preliminary level
Equipment data sheets for major equipment to preliminary level
Specifications accompanying enquiry packages returned from service providers.
Battery limits and technical specifications for use with designed packages
Preliminary overall plot plan drawing
Preliminary overall site layout drawing
Updated process description
Updated Process Design Criteria (PDC)
Updated Block Flow Diagram (BFD)
Process Flow Diagrams (PFDs)

21.2.2.7. Pricing Basis

Pricing for the direct works for the Prefeasibility study cost estimate is based on the following sources:

Mechanical Equipment List
Formal budgetary and informal email quotations for various design packages
Quotations from specialist turnkey suppliers for various mechanical equipment packages
In-house database information
Factored cost estimation from the mechanical equipment cost
Estimated allowances from previous project experiences and norms.

21.2.2.8. Presentation of Capital Cost

The overall capital cost estimate was compiled using Excel spreadsheet format.

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21.2.2.9. Capital Estimate Structure

The cost estimate is compiled in line with the scope of works and formatted around the SGS Bateman estimating best practice guide the traditional code of account for direct costs, indirect costs and EPCM costs calculation.

21.2.2.10. Direct Field Costs (DFC)

The direct field costs include permanent materials and equipment, freight to site, construction labour and equipment including Preliminary and General costs associated with contractors' supervision, overheads and profits, temporary construction facilities, construction mobile equipment, accommodation of construction labour, and contractor mobilisation and demobilisation.

21.2.2.10.1. Earthworks

The value of the earthworks is calculated as a percentage factor of the mechanical equipment supply cost using factors and norms from past projects that are of a similar nature. Preliminary and general costs of the earthworks are included in the capital cost as a part of the factored percentage of the Civils Works P&Gs' estimate cost estimate and are indicated separately.

Contractor's indirect costs (P&G's) cater for the contractor's mobilisation and demobilisation including establishment and later removal of construction plant and equipment, contractor's manual indirect and non-productive labour, scaffolding, safety equipment, personal protective equipment, transport and travelling, accommodation and supervision including contractual requirements relating to finance costs, insurance, bonds, and work permits.

21.2.2.10.2. Civil Works

The value of civil works is calculated as a percentage factor of the mechanical equipment supply cost using factors and norms from past projects that are of a similar nature. Preliminary and general costs are included in the capital cost estimate as a part of the factored percentage of the Civils Works P&Gs' estimate cost and will be indicated separately.

It has been assumed that ground conditions are suitable for standard concrete equipment support structures, with no piling or special foundations requirements.

21.2.2.10.3. Buildings and Infrastructure

The value of the buildings and infrastructure costs have been estimated by CREO Engineers.

21.2.2.10.4. Structural Steel

The value of structural steelwork supply has been calculated as a percentage factor of the mechanical equipment supply cost using factors and norms from past projects that are of a similar nature. The installation cost is priced using a factored percentage of the supply cost based on factors and norms from previously executed projects. Preliminary and general costs of the structural steelwork erection are included in the capital cost estimate as a part of the factored percentage of the SMPP P&Gs' estimate cost and are indicated separately.

21.2.2.10.5. Platework and Liners

The value of platework, including tanks, launders, hoppers, chutes and sump costs has been calculated using the quantities provided on the latest issued equipment lists and/or schedules of equipment. Supply rates used to calculate the costs have been based on historical database information from recently completed SMPP contracts. The erection cost has been calculated by using typical similar project rates of

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recently completed projects. Preliminary and general costs of the platework installation are the contractors' indirect costs and are included in the capital cost estimate as a part of the factored percentage of the SMPP P&Gs' estimate cost and indicated separately.

The platework and associated linings cost include supply, shop detailing, fabrication, surface protection (where applicable), freight and installation of all shop and site-fabricated platework and associated linings. Rubber lining, epoxy internal surface treatment and liner plate costs included where required.

21.2.2.10.6. Mechanical Equipment

The value of mechanical equipment supply cost has been based on supply prices for the items identified in the mechanical equipment list developed from the PFD's and mechanical layouts drawings. The installation cost has been priced using a combination of quoted installation costs and factored percentage of the supply cost based on factors and norms from previously executed projects. Preliminary and general costs of the mechanical equipment installation are included in the capital cost estimate and indicated separately as a part of the factored percentage of the SMPP P&Gs' estimate cost.

Enquiries were issued to multiple vendors or to single source vendors where specialist equipment suppliers have been motivated in a sole source motivation.

A mixture of full formal quotations, budget quotations and in-house database information has been used and escalated to the current cost estimate base date as defined in the project procurement strategy and procurement operating plan.

Equipment installation has been calculated by using a factored percentage of equipment supply cost. Preliminary and general costs have been calculated by using a factored percentage of the erection cost and included in the capital cost estimate as a part of the factored percentage of the SMPP costs which indicated separately.

21.2.2.10.7. Piping and Valves

The value of the piping and valves supply has been calculated as a percentage factor of the mechanical equipment supply cost using factors and norms from past projects that are of a similar nature, apart from the Kiln Scrubber Plant which has been included in the supply cost from the equipment supplier. The installation has been priced using a factored percentage of the supply cost based on factors and norms from previously executed projects. Preliminary and general costs for the piping and valves installation are included in the capital cost estimate and has been indicated separately as a part of the factored percentage of the SMPP P&G's estimate cost.

21.2.2.10.8. Electrical

The value of the electrical supply has been calculated as a percentage factor of the mechanical equipment supply cost using factors and norms from past projects that are of a similar nature. The installation cost has been priced using a factored percentage of the supply cost based on factors and norms from previously executed projects. Preliminary and general costs for the piping and valves installation are included in the capital cost estimate and has been indicated separately as a part of the factored percentage of the E&I P&Gs' estimate cost.

21.2.2.10.9. Control and Instrumentation

The value of the control and instrumentation equipment supply has been calculated as a percentage factor of the mechanical equipment supply cost using factors and norms from past projects that are of a similar nature. The installation cost has been priced using a factored percentage of the supply cost based on factors and norms from previously executed projects. Preliminary and general costs for the piping and valves installation are included in the capital cost estimate and has been indicated separately as a part of the factored percentage of the E&I P&Gs' estimate cost.

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21.2.2.10.10. Preliminary and General

Preliminary and general costs are the erection contractors indirect costs; which make provision for the contractor's mobilisation and demobilisation including establishment and later removal of construction plant and equipment, contractor's manual indirect and non-productive labour, scaffolding, maintaining health and safety, personal protective equipment, transport and travelling, accommodation and supervision including contractual requirements relating to finance costs, insurance, bonds and work permits.

21.2.2.10.11. Allowances

Growth allowance is applied to estimates where materials have been quantified, and where part of the estimated cost is done in more detail, then the allowance for growth will be agreed in consultation with the discipline engineers taking into consideration the growth allowances recommended by SGS Bateman Estimator's Best Practice Guide PCNG-0920-002 Rev 1 guideline for those specific portions of the estimate.

A percentage allowance for factors is applied to the estimate based on the degree of engineering completed and quality of pricing information that supports the estimate.

Productivity allowances are dependent upon the various areas that are going to be worked in and may require in-house research according to the experience of working within a particular country. Productivity allowances were not utilized in this Capex as factorized percentage allowances were used to calculate the installation costs.

Allowance for design growth and wastage were agreed in consultation with the discipline engineers taking into consideration the growth allowances recommended by SGS Bateman Estimator's Best Practice Guide PCNG-0920-002 Rev 1 guideline for a Class 3 estimate classification.

21.2.2.10.12. Transportation and Logistics

The costs for logistics, sea freight and inland transportation has been included in the estimate for the delivery of equipment and materials from the country of supply to the site. This cost has been estimated including import duties where necessary based on the procured market quotations or based on information and norms based on recently completed projects.

21.2.2.10.13. Spares

Spares and consumables have been handled as follows at this stage:

Commissioning and Initial Spares are included in the capital estimate as a factored percentage of the mechanical equipment cost.
Critical spares are included as a percentage factor in the capital estimate as a factored percentage of the mechanical equipment cost.
Operational Spares are excluded from the capital estimate but included in the Opex estimate.

21.2.2.10.14. First Fills (Oils, Lubricants)

Cost for first fills of lubricants is included as a factored percentage of the mechanical equipment for the packages which the vendors did not have the cost included in the received budget quotations.

21.2.2.10.15. Vendor Assistance

Cost for vendors assistance during construction and commissioning is included as a factored percentage of the mechanical equipment for those packages which the vendors did not have the cost included in the received budget quotations.

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21.2.2.11. Indirect Field Costs (IFC)

Indirect costs are generally duration based and include items that are necessary for the completion of the project that are not related to the direct construction costs.

The project indirect field costs are dependent on the project duration.

21.2.2.11.1. Engineering, Design and Project Management

Project management, engineering and procurement costs have been included as a percentage allowance of the total DFC and this cost covers the construction management and commissioning assistance up to C2 (EPCM) and costs directly associated with the implementation of the project. No Allowance has been made for training of local personnel and defect liability period.

21.2.2.11.2. Bonds, Guarantees etc.

An allowance for bonds and guarantees based on a factored percentage is included in the estimate.

21.2.2.11.3. Project Insurance

Insurance included in the estimate is an allowance for project related risks which are insurable. It is dependent on project variables and project specific circumstances. It typically includes the following:

Contractors All Risk on construction and site activities typical cover – this depends on the extent of cover required.
Third Party Liability insurance typical cover.
Medical Evacuation and CasEvac typical cover. This depends on the area, location, and the detailed circumstances.
Marine Cargo and difference in excess typical cover.

The following risks were not allowed for in the estimate and thus excluded due to the specific requirements the owner may have. These should be strongly considered in addition to those listed above:

Delay in Start-up insurance (DSU).
Project Specific required professional indemnity.
Advance Loss of Profits (ALOP).

The insurance estimate allowance should be finalized by performance of an insurance review by specialized parties, once the exact requirements of the owner and the project are available in more detail. An allowance based on a factored percentage of the total net cost is included in the estimate as a guideline. However, it is advisable that formal quotations for project insurance are sourced and issued by the client.

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21.2.2.11.4. Project Contingency

Project Contingency is shown for clarity in this section. A more detailed discussion is available in Section 22.

21.2.2.12. Owner's Cost

The estimate specifically excludes Owner's costs.

Generally, owner's costs are, but not necessarily limited to, the following:

Owner's project management team and consultants
Pre-production, commissioning costs, etc
Provisions for fluctuations in ROE and escalation
Owner's contingency for changes in scope or additional work
Pre-development costs (cost of study, etc.)
Land acquisition
Resettlement or relocation costs
Community relations
Business systems
Loss of production and efficiency resulting from implementation
Owner's start-up and commissioning crew
Project taxes, fees, duties, customs, permitting and approvals.
Development fees and approval costs of statutory authorities
Finance fees or cost of capital
Pre-production costs (operator training)
Workplace health and safety fees
Operational readiness
Site survey & soils testing
Environmental considerations (EIA)
Additional study fees

21.2.2.13. Exclusions

The following items are excluded from the scope of this estimate:

Costs of geotechnical investigations, site surveys, metallurgical test work, EIA permitting and regulatory compliance requirements, together with any costs associated with process design changes necessitated by the outcome of such investigations.
Any costs associated with statutory requirements, local permits, licensing, royalties and approvals, social, community or environmental requirements.
Owner's costs
Owner's contingency allowance
Value Added Tax (VAT) and Goods & Services Tax (GST)
Financing costs
Financial modelling
Marketing analysis
Business system costs
Pre-development and operational readiness costs
Operational costs (included in the Operating Costs Estimate)
External auditing costs
Permit applications

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Foreign currency exchange rates variations from the estimate base date
Schedule acceleration costs
Schedule delays and associated costs, such as those caused by:
Unexpected site conditions
Weather conditions other than fair
Unidentified ground conditions
Labour disputes
Force Majeure
Testwork and laboratory costs unless specifically included.
Provision of landscaping and nursery services
Closure/rehabilitation costs
Higher level management system (MIS, MES or ERP)
Infrastructure and facilities outside process plant fence
Topographical surveys and geotechnical investigations
Hydro-geological and hydrological evaluation
Electrical power supply to the site distribution substation
Process and portable water supply to the site bulk water storage tanks
Environmental impact assessment, regulatory permitting and/ or licencing, and related studies
Work relating to mining support infrastructure.
Facilities outside of the process plant battery limits, (e.g., external access roads and haul roads)
Work relating to socio-economic considerations.
Tailings storage facility
Translation services and associated costs
Additional work that might be required but that is not specifically included in the above-mentioned scope of work.

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Table 133 Capital Estimate Summary (USD)

Description	A Supply Cost	% Total A	B Erection Cost	Total Cost A+B	% Total DFC
Direct Field Costs
Bulk Earthworks		0.00%	$4 607 414	$4 607 414	3.30%
Infrastructure (Excluded - CREO)		0.00%			0.00%
Civil Works		0.00%	$4 465 647	$4 465 647	3.20%
Buildings Architectural (Excluded - CREO)		0.00%			0.00%
Structural Steelwork	$8 371 715	9.30%	$1 506 909	$9 878 623	7.00%
Platework & Liners	$3 712 757	4.10%	$402 047	$4 114 803	2.90%
Conveyor Mechanicals	$1 034 532	1.10%	$341 395	$1 375 927	1.00%
Mechanical Equipment	$51 064 142	56.80%	$6 126 656	$57 190 798	40.60%
Piping & Valves	$6 250 880	6.90%	$3 437 984	$9 688 864	6.90%
Overland Piping - Excluded
Electrical	$8 371 715	9.30%	$1 674 343	$10 046 058	7.10%
Instrumentation	$5 581 143	6.20%	$1 116 229	$6 697 372	4.80%

E/W & Civil Works P&G's			$4 536 530	$4 536 530	3.20%
SMPP P&G's		0.00%	$15 862 551	$15 862 551	11.30%
E&I P&G's		0.00%	$6 697 372	$6 697 372	4.80%

Transportation of Equipment (DAP) to site	$3 848 641	4.30%		$3 848 641	2.70%
Commissioning Spares	$651 233	0.70%		$651 233	0.50%
First fill of lubricants - Excl First Fill of Chemical Reagents	$299 475	0.30%		$299 475	0.20%
First fill of Chemical Reagents - Excluded					0.00%
Vendor assist during Constr & Comm	$781 480	0.90%		$781 480	0.60%

TOTAL DIRECT FIELD COSTS	$89 967 712	100.00%	$50 775 075	$140 742 788	100.00%
Home Office & Indirect Field Costs
EPCM (Excl Genet) @ 15% of DFC Costs	$21 111 418	23.50%		$21 111 418	15.00%
EPCM - Genet Front End @ 5% of DFC Costs	$782 924	0.90%		$782 924	0.60%

TOTAL H.O. & INDIRECT FIELD COSTS	$21 894 342	24.30%		$21 894 342	15.60%
TOTAL NET COST	$111 862 054		$50 775 075	$162 637 129

Other Costs					% of TPC
Bonds Guarantees etc. @ 0.25% of TNC	$406 593	0.50%		$406 593	0.20%
Insurance	$2 869 570	3.20%		$2 869 570	1.30%
Contingency	$41 478 323	46.10%		$41 478 323	18.60%

TOTAL OTHER COSTS	$44 754 485	49.30%		$44 754 485	20.10%
Owner's Costs - Excluded
Genet front end	$15 071 855	16.80%	$586 620	$15 658 475	11.10%

TOTAL OWNER'S COST	$15 071 855	16.80%	$586 620	$15 658 475	11.10%
OVERALL PROJECT COST	$171 688 394		$51 361 695	$223 050 090
Overall Project Cost Without Contingency	$130 210 071		$51 361 695	$181 571 767

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21.2.3. Tailings Capital Costs

Tailings capital is estimated based on the requirements of the mine and process plan, with the initial berms placed during the pre-production period, and a further expansion in Year 2 of operations. For cost model scheduling purposes, Initial tailings berm construction occurs in Year -1. Contingency on the tailings construction is applied in the year following construction. Overall Project Contingency will be addressed in Section 22.

Table 134 Tailings Capital Costs (USD)

Description	Total LOM Capital	Initial Capital Cost (Before start of Deposition)	Sustaining Capital Cost (Yr 2 - 20)
		USD	USD
PRELIMINARY AND GENERAL	$3 016 335	$2 459 566	$ 556 769
EARTHWORKS	$4 672 986	$3 363 477	$1 309 509
GEOSYNTHETICS	$11 407 135	$4 423 927	$6 983 208
PIPEWORKS AND APPURTENANCES	$ 625 175	$ 137 153	$ 488 021
GEOTECHNICAL INSTRUMENTATION	$ 276 947	$ 58 256	$ 218 691
INFRASTRUCTURE	$ 270 016	$ 215 749	$ 54 266
ENGINEERING	$1 321 658	$ 409 931	$ 911 727
SUB TOTAL	$21 590 251	$11 068 059	$10 522 192
CONTINGENCY	$3 924 430	$1 617 039	$2 307 391
Total With Contingency	$25 514 681	$12 685 098	$12 829 583

21.2.4. Non-Process Infrastructure

Non-Process Infrastructure includes the capital costs for facilities listed in Section 18. The services and infrastructure covered include:

Site Wide Services & Infrastructure
Bulk Water Supply
Bulk Power Supply
Electrical Distribution
IT / Communications Infrastructure
Site Access & Internal Roads
Plant Fencing
General Facilities & Buildings
Supplementary E&I
Plant Buildings
Transportation & Site Vehicles
EPCM for Non-Process Infrastructure

Table 135 below provides a breakdown of the estimated infrastructure capital.

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Table 135 Non-Process Infrastructure Capital Costs (USD)

Direct Field Cost	$54 394 984
General Facilities & Buildings	$7 929 006
Plant Buildings	$7 339 306
Ablutions	$35 375
Change Houses	$224 463
Engineering Workshop	$986 651
Front End Office	$40 429
Front End Workshop	$77 495
Fuel Farm	$103 122
General Admin & Engineering Services	$1 447 015
Plant Buildings	$2 632 357
Plant Sewerage Treatment	$289 440
Reagent Storage	$631 577
Sewerage Distribution	$37 300
SHE, Training & Security	$70 750
Warehousing & Stores	$680 789
Wash Bays	$82 542
Transportation & Site Vehicles	$589 701
Site LDV's	$589 701
Site Wide Services & Infrastructure	$46 465 977
Bulk Power Supply	$29 884 283
Grid Supply	$29 884 283
Bulk Water Supply	$9 732 708
Borehole / Main Bulk Water Pipeline	$6 945 240
Borehole Pump / Reservoir Installations	$619 274
Electrical and I&C	$748 541
Fire Water Storage	$130 324
Potable Water Supply	$237 271
Raw Water Storage	$1 052 058
Electrical Distribution	$733 403
Cable Racking and Trenching	$122 240
Earthing and Lightning Protection	$1 522
Electrical Distribution	$78 230
General Site Lighting & Layout	$160 990
Internal LV distribution for Admin (Kiosks & Distribution)	$247 734
Internal MV Distribution	$122 687
ICT Infrastructure	$56 005
Onsite Communications	$56 005
Plant Fencing	$94 093
Site and Specific Area Fencing	$94 093
Site Access & Internal Roads	$5 965 485
Internal Site Roads	$218 632
Main Site Access Road	$5 516 551
Site Parking & Main Access (Visitor, LDV & Busses)	$130 600
Weighbridge	$99 702
Indirect Field Costs	$4 351 599
EPCM	$4 351 599
Total Non Process Infrastructure without Contingency	$58 746 582
Contingency	$11 749 316
Total Non Process with Contingency	$70 495 899

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21.3. Project Operating Costs

The overall operating costs for mining, processing, Site services and G&A are shown in Table 136.

The mining cost per tonne for all tonnes averages $2.63 over the life of mine. The variance between mined ore tonnes and processed ore tonnes reflects the reduction of gangue tonnage achieved through ore sorting.

The processing summary in Table 136 also includes concentrate transport costs, and tailings costs.

Table 136 Operating Cost Summary (USD)

Area	LOM Total Cost	Cost Per Ore Tonne Mined	Cost Per Ore Tonne Processed	Cost Per kg TREO Recovered (USD/t)
Mining	$652 439 644	$20.38	$37.32	$24.78
Processing	$996 304 067	$31.13	$56.98	$37.84
General and Administration	$29 783 297	$0.93	$1.70	$1.13
Total Operating Cost	$1 678 527 008	$52.44	$96.00	$63.75

21.3.1. Process Plant Operating Cost Estimate

The following assumptions have been made in estimating the operating costs:

A mass balance was completed for the circuit to anticipate mass flow through the process, using information derived from the testwork latest results. Reagent consumption and required supply has been derived from this balance.
Preliminary equipment sizing was evaluated from received vendor budget quotes such as to determine motor sizing. Motor sizing was calculated where supplier estimates were not available.
Reagent pricing was obtained from Protea Chemicals and other specialist vendors including Florrea. Reagent consumptions were determined via mass balance and based on the data found in the Process Design Criteria and testwork
Acid pricing supplied by Namibia Critical Metals.
Maintenance costs were estimated as a factor of the mechanical equipment installed costs. This was estimated from the mechanical equipment capex estimation as per SGS Bateman standards and referenced to the Mintek Handbook (March 2002).
Power estimations from the basic equipment sizing were costed using the supplied power tariff from CREO.
Namibian Labour rates provided by Namibia Critical Metals.
Costs were determined for typical operating duty at design throughput.
No contingencies have been included in the operating cost estimates unless otherwise stated.
Costs are expressed in United States Dollar (USD).
All costs are in 2025 terms, unless otherwise specified.

The operating cost estimate, based on nameplate mill feed, is summarised in Table 137 and shown in Figure 202. The estimate includes reagents and consumables, fuel, labour, maintenance materials and power consumption.

Final process costs use the nameplate values per tonne, applied to the mill feed profile. Extra costs associated with processing not included in Table 137 are concentrate transport at $42.15 per tonne of concentrate, and tailings handling costs of $0.22 per tonne.

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Table 137 Operating Cost Estimate Nameplate Capacity

Cost Component	(USD/a)	%
Labour	6 074 413	7.2
Maintenance	5 663 124	6.7
Power	12 585 404	15.0
Fuel (Genet Equipment)	936 114	1.1
Genet (operating cost for front end)	9 420 309	11.2
Reagents and Consumables	49 342 167	58.7
TOTAL	84 021 532	100

Feed (t ROM/a)	3 010 000
$/t of ROM Feed	27.91
Feed to mill (t/annum)	1 482 000
$/t of mill Feed	56.7
$/kg Product	36.82

Figure 202 Opex Distribution

21.3.1.1. Scope

The operating costs can be categorised as fixed or variable costs.

Fixed costs include:
- Manpower (Labour for plant operation and maintenance only)
- Maintenance and operating supplies

Variable costs include:
- Power
- Reagents, fuel and consumables
- Waste Handling

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21.3.1.2. Accuracy

The accuracy for the operating cost estimate is as for the capital cost estimate. The methodology used in preparing the Opex costs in this report was based on the scope, pricing and information available at the time but should be revised as the industry fluctuates, especially in terms of reagent pricing. During commissioning, start-up and ramp-up unit costs will vary in comparison to when the plant is operating at full capacity.

21.3.1.3. Exclusions

The following are excluded from this process plant operating cost estimate:

Mining Costs
Tailings Handling Costs
General, Medical and Administration Costs, other than plant and technical/engineering services
Security costs
Duties and taxes on exports of products
Marketing costs
Depreciation and replacement capital
Insurance
In-country corporation tax
First fill reagents costs (included in capital estimate)
No provision for annual increases in salary, services and supplies growth has been allowed
Contingency

21.3.2. Fixed Costs

21.3.2.1. Labour

Labour was costed and detailed on the project using supplied rates from NCM (17 July 2024)

The following assumptions have applied to the labour cost estimation:

2% escalation to September 2025 base date
The type and level of personnel required has been based on estimates for skilled, unskilled, technical, engineering and senior engineering personnel but would require further review by the Namibia Critical Metals.
The monthly all-inclusive cost to company of personnel was provided by the Namibia Critical Metals.
Includes production, maintenance and SHEQ personnel only. Assumes Medical, Admin, Finance and Senior HR personnel are included in overheads.
Manning requirements based on an 8 hourly shift rotation and typical SA process plant case to meet SA regulations in terms of health and safety act.

The labour requirements are seen below:

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Table 138 Process Labour Requirements
| | Labor complement per shift | | | | | Total cost per month | Total cost per year |
| --- | --- | --- | --- | --- | --- | --- | --- |
| | Morning | Afternoon | Night | 4th Shift | Total | | |
| Metallurgical Operation Total (All Shifts) | 7 | 7 | 7 | 7 | 28 | $45 018.83 | $540 226.01 |
| Hydrometallurgical Operation Total (All Shifts) | 9 | 9 | 9 | 9 | 36 | $61 159.37 | $733 912.40 |
| Reagents and Utilities Total (All Shifts) | 14 | 14 | 4 | 14 | 46 | $69 874.21 | $838 490.56 |
| Process Plant Technical Total (All Shifts) | 48 | 16 | 13 | 10 | 87 | $330 148.64 | $3 961 783.62 |
| Total | 78 | 46 | 33 | 40 | 197 | $506 201.05 | $6 074 412.59 |
| Plant Throughput (t/annum) | | | | | | | 3 010 000 |
| Operating Cost ($/t ROM) | | | | | | | $2.02 |

21.3.2.2. Maintenance

In all cases, the cost of maintenance supplies is calculated as a factor of the mechanical equipment supply costs excluding piping and valves, electrical and instrumentation based on previous studies for typical uranium hydrometallurgical plants but excludes the maintenance associated with the acid plant. Generally, maintenance materials are considered to be (7 - 13%) of the mechanical equipment supply cost. An estimate of 10% has been applied. The maintenance labour component has been allowed for in the annual labour estimate.

21.3.3. Variable Costs

21.3.3.1. Power

The power cost estimate is based on the supplied power tariff structure from infrastructure at a unit price of 134 USD/MWh. This cost is assumed to include: -

Service and Administration charges
Network Capacity charges
Network Demand charges
Ancillary service charges
Energy charges
Any other charges that are routinely levied to the client.

Operating power cost for continuous standby power generation (including monthly test-runs) are not included in this cost.

For the proposed flowsheet, basic equipment sizing was completed to determine Absorbed Power (kW) for each unit operation in the process. Where installed power was used a power factor of 0.80 was largely applied.

21.3.3.2. Reagents and Consumables

Reagent consumptions were determined from testwork and input into the process mass balance. The completed mass balance thus allowed for a reagent consumption calculation on a t/h basis.

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Reagent prices were obtained from local suppliers as best available. With reagent supply pricing from Protea Chemicals Namibia and Florrea used in the Opex estimate. Acid costs supplied by preliminary pricing from Namibia Critical Metals.

Only major consumables considered. Minor consumables assumed to be 5% of overall annual cost. Minor consumables to include drums, packaging, minor chemicals, office consumables, lab consumables

Crusher liner wear excluded and allowed for in maintenance costing.

All reagent pricing is in September 2025 terms. Reagent transport costs estimated based on haulage costs per km for payloads provided by CREO Engineers from Walvis Bay to Lofdal site.

Table 139 Reagent Costing

Reagent and Consumables	Specific Consumption		Annual Consumption	Annual Cost
	g/t mill feed	kg/t hydro feed	t/a (litres/annum for fuel)	USD/a	USD/t
Sodium Silicate (N Type)	213		315.67	149 310.02	473
Calgon	46		68.17	166 066.99	2 436
Florrea 3900	1 235		1 830.27	12 537 349.50	6 850
Florrea 3000	260		385.32	4 045 860.00	10 500
NaOH	100		148.20	114 992.83	775.93
H_{2}SO_{4}	25 749	757	38 160.00	5 914 799.26	155
MgCO_{3}	25 466	749	37 740.61	11 963 774.00	317
Na_{2}CO_{3}	1 168	34.36	1 731.33	857 009.18	495
Magnafloc 10 (tails thick)	40		59.28	138 122.40	2 330
Magnafloc (conc thick)	1	0.04	2.02	4 696.16	2 330
Coagulant (tails thick)	70		103.74	185 711.20	1 790.16
Coagulant (conc thick)	2	0.07	3.53	6 314.18	1 790.16
Ca(OH)_{2}
NH_{4}OH	9	0.25	12.60	7 148.17	567.45
H_{2}O_{2}	31	0.9	45.35	53 965.55	1 190
Ancillaries (% of Reagent Cost)				1 807 255.97
Grinding Media				4 561 702.61
Raw Water			1 482 192.00	508 391.86
Fuel - acid bake kiln			7 234 926.05	6 319 697.48
Total Consumables cost				49 342 167.36

21.3.3.3. Fuel

Diesel requirements and costing has been included for the running of the Genet ore handling vehicles. Fuel costs were included for the operation of the acid bake kiln. A diesel price of 15.92 NAD/L was used.

Equipment used for the purposes of maintenance are not included as this is covered in the total maintenance costing.

Transportation within plant battery limits not considered. Any external travel requirements beyond plant boundaries not considered.

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21.3.3.4. Power and Water Cost

The main energy source for the various mill site process plant operations is electrical power. The power cost is a combination of grid power supplied by NamPower and solar power provided by Agrekko.

An allowance was made for the cost to purchase raw water from the local water authorities. This is indicated in the Table 21-11.

Table 140 Mill Site Process Plant Energy and Water Cost

Description	Unit	Unit Cost (USD)	Price Basis	Information Source
Power Price	kWh	$0.134	Namibia	NamPower / Agrekko
Raw Water	m³	$0.343	Namibia	Estimate

21.3.3.5. Site Closure

Site closure costs are related to the progressive reclamation of the tailings facilities, with overall sustaining costs of $1 039 987 over the life of mine.

For this PFS, no costs are provided for final site closure and reclamation.

21.4. Mine Operating Costs

The mine operating costs are derived from contractor quotes based on the PFS mine plan, as well as first principle costs for mine indirects.

The components of the mine operating costs include:

Contractor Equipment and labour for loading, hauling, drilling and blasting
Contractor fixed plant costs
Vendor quotes for fuel, explosives and support equipment
Factors for maintenance, ground engaging tools, lubricants and tire consumption

Indirect mining costs are based on:

Equipment and labour for pit support equipment
Contractor equipment rental rates
Contractor capital recovery on the rentals
Fuel Consumption based on factored hours for indirect equipment
Factors for maintenance, lubricants and tire consumption

Haulage rates increase with pit depth and haul distance based on the mine plan. Drill and blast costs include pre-split drill and blast.

Table 141 provides an annual breakdown of mining costs by key area.

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Table 141 Mine Operating Cost Summary

Mining Operating Costs		Totals	Pre-Production (costs under Capital)	Y1	Y2	Y3	Y4	Y5	Y6	Y7	Y8	Y9	Y10	Y11	Y12	Y13
High Grade Ore	t	12 643 731	341 560	681 340	797 308	1 203 392	1 100 000	881 800	1 276 297	726 153	1 230 922	1 008 982	828 168	843 920	1 139 446	584 442
Low Grade Ore	t	19 362 902	259 167	774 559	1 081 119	1 391 155	1 528 000	1 594 064	1 688 775	1 092 378	1 562 783	1 854 244	2 408 882	1 097 605	2 592 622	437 548
Waste Mined	t	205 242 156	4 070 257	10,845,291	16,415,922	18,823,331	20,032,606	20,077,128	19,854,222	20,787,038	19,837,259	19,845,790	15,229,929	12,692,712	6,518,619	212,052
Marginal Ore to Stockpile	t	10 904 963	129 017	398 809	605 651	727 055	1 062 394	970 008	703 706	717 431	864 036	1 113 984	1 533 021	865 762	1 084 828	129 260
Total Material Mined	t	248 153 751	4 800 000	12 700 000	18 900 000	22 144 934	23 723 000	23 523 000	23 523 000	23 323 000	23 495 000	23 823 000	20 000 000	15 500 000	11 335 515	1 363 302
Low Grade Ore to Ore Sorter	t	19 362 365		726 400	1 284 000	1 495 600	1 528 000	1 528 000	1 528 000	1 191 000	1 191 000	2 191 472	2 324 366	1 189 099	2 367 288	818 140
Sorted Ore to Mill	t	4 840 591		181 600	321 000	373 900	382 000	382 000	382 000	297 750	297 750	547 868	581 091	297 275	591 822	204 535
High Grade Ore to Mill	t	12 643 731		522 900	924 200	1 076 500	1 100 000	1 100 000	1 100 000	1 184 250	1 184 250	934 132	900 909	892 702	890 178	833 710
Total Mill Feed	t	17 484 322		704 500	1 245 200	1 450 400	1 482 000	1 482 000	1 482 000	1 482 000	1 482 000	1 482 000	1 482 000	1 189 977	1 482 000	1 038 245
Strip Ratio (Marginal Ore Included with Waste)		6.75		7.72	9.06	7.54	8.03	8.50	6.93	11.83	7.41	7.32	5.18	6.98	2.04	0.33
Operating Costs
Drilling		$67 883 137		$3 628 707	$5 118 691	$6 062 765	$6 518 093	$6 400 458	$6 582 684	$6 044 170	$6 415 996	$6 515 932	$5 754 628	$4 266 299	$3 909 014	$665 699
Blasting		$73 738 405		$3 826 844	$5 668 869	$6 685 115	$7 149 244	$7 077 577	$7 124 237	$6 954 453	$7 096 853	$7 198 814	$6 122 434	$4 691 216	$3 649 587	$493 162
Loading		$89 690 934		$5 028 549	$6 986 892	$8 067 499	$8 816 746	$8 681 672	$8 779 914	$8 312 751	$8 487 894	$8 609 945	$7 328 785	$5 611 388	$4 382 663	$596 234
Hauling		$292 085 997		$11 769 003	$18 260 483	$22 276 718	$25 909 325	$25 600 759	$24 974 370	$27 013 162	$28 679 853	$35 112 316	$25 689 282	$24 714 578	$19 273 324	$2 812 824
Secondary Support, Rehandle and Other Mining		$35 740 547		$1 716 508	$2 273 575	$2 875 466	$2 355 449	$2 689 048	$2 343 944	$4 484 229	$3 414 235	$4 673 163	$3 902 824	$2 510 180	$1 478 825	$1 023 101
Contractor Fixed Plant Cost		$64 671 197		$4 677 598	$4 705 701	$4 667 389	$4 683 543	$4 687 445	$4 687 445	$4 687 445	$4 687 445	$4 687 445	$4 687 445	$4 687 445	$4 687 445	$8 437 402
Mining Indirects (Owner's Costs)		$28 629 428		$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 157 947	$2 734 061
Total Mining Opex		$652 439 644		$32 805 157	$45 172 159	$52 792 899	$57 590 348	$57 294 907	$56 650 542	$59 654 158	$60 940 224	$68 955 562	$55 643 345	$48 639 054	$39 538 806	$16 762 483
Mining Cost Per Tonne		$2.63		$2.58	$2.39	$2.38	$2.43	$2.44	$2.41	$2.56	$2.59	$2.89	$2.78	$3.14	$3.49	$12.30

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Secondary Support, Rehandling and Other Mining Costs include:
- Secondary & Support Mining
- ROM Rehandling
- Stockpile Rehandling
- Surface Haul Road Construction
- Surface Haul Road Maintenance
- Clear, Remove and Grub Pit & Waste Rock Dump Areas
- Topsoil Loading, Haulage and Stockpiling

Contractor fixed plant costs include contractor pricing for the operation of their fixed plant equipment and service vehicles.

Mine Indirects include costs for:
- LDV Fuel & Maintenance
- Computers & Hardware
- Software License Fees (Mine Planning, Geology & Survey)
- Training & Membership Fees
- Special Projects
- PPE, General & Office Consumables
- Grade Control & Resource In-fill Drilling
- Owner Mining Labour
- Mining Manager
- Superintendent Technical Services (Chief Geologist)
- Grade Control Geologist
- Production Control Geologist
- Geology Field Assistant
- Resource Geologist
- Geotechnical Engineer
- Geotechnical Field Assistant
- Senior Mine Surveyor
- Mine Surveyor
- Surveyor Assistant
- Draftsman
- Superintendent Mining (Chief Mining Engineer)
- Short Term Planning Engineer
- Medium / Long Term Planning Engineer

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21.5. General and Administration

General and Administration costs include costs for site administration, and site services.

G&A cost components include:
- Administration Labour
- Annual Allowances for office supplies
- Annual Allowances for Environmental sampling
- Non-Process water use
- Non-Process site facility Power
- Annual IT / Network Support
- Site material transport / deliveries

Annual site G&A is summarized in Table 142

Table 142 Annual General and Administration Costs
| Personnel | $1 254 086 |
| --- | --- |
| Office Supplies | $ 150 000 |
| Environmental Compliance | $ 160 000 |
| Non-Process Water Usage | $ 20 930 |
| Non-Process Site Facility Power | $ 195 312 |
| Internal Road Maintenance (Non Mining) | $ 44 552 |
| IT / Network Support | $ 2 001 |
| Site Transport and Deliveries | $ 464 144 |
| Total Annual G&A | $2 291 023 |

Administration labour includes the positions in Table 143

Table 143 Administration Labour
| Administration | Quantity |
| --- | --- |
| NCM General Manager | 1 |
| Manager Administration | 1 |
| HR Superintendent | 1 |
| HR Assistant | 1 |
| Finance Superintendent | 1 |
| Accounts Payable | 1 |
| Accounts Receivable | 1 |
| Purchaser | 1 |
| Warehouse Supervisor | 1 |
| Shipper / receiver | 2 |
| Warehouse Floor Staff | 4 |
| Manager HSE | 1 |
| SHEQ Officer | 1 |
| Training Officer | 1 |
| Environment Compliance Officers | 1 |
| Environment Technicians | 2 |
| First Aid / Medical Staff | 4 |
| Fire / Mine Rescue Response | 4 |
| Total G&A Personnel | 29 |

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22. Economic Analysis

22.1. Introduction

The economic evaluation of the Lofdal Rare Earths Project as presented in this PFS assumes the project will be 100% equity financed.

This economic evaluation uses parameters relevant as of September 2025, under conditions likely to be applicable to project development and operation and analyses the sensitivity of the project to changes in the key project parameters.

Mining and treatment data, capital cost estimates, operating cost estimates and rare earth oxide prices have been put into a financial model to calculate the internal rate of return (IRR) and net present value (NPV) based on calculated project after tax cash flows.

The current market exhibits several different pricing levels for individual rare earth elements. As such, the financial model is calculated for two cases; the Base Case using export controlled Chinese pricing and the Divergent Price model.

The scope of the financial model has been restricted to the project level and as such, the effects of interest charges and financing have been excluded.

The model includes sensitivity analyses to demonstrate the effect of variations in key parameters on the economic returns from the project.

Estimated project returns and the key financial statistics are summarized and discussed in this section and are supported by tables and charts. A summary of the financial model results for the project is included in Table 144.

Table 144 Summary of Financial Analysis

Metric	Base Case	Divergent Pricing
Net Present Value (NPV, discount rate 5%)	Pre-tax: USD389.2 million
After-tax: USD275.5 million	Pre-tax: USD1,245.6 million
After-tax: USD747.9 million
Internal Rate of Return (IRR)	Pre-tax: 21.7%
After-tax: 19.0%	Pre-tax: 44.1%
After-tax: 34.8%
Life-of-Mine Nominal Cash Flow	Pre-tax: USD709.6 million
After-tax: USD513.1 million	Pre-tax: USD2,027.4 million
After-tax: USD1,242.3 million
Pre-Production Capital Costs	USD273.4 million	Same as Base Case
Total Capital Costs	USD347.9 million (including contingency of USD57.4 million)	Same as Base Case
Capital Payback Period (after-tax)	4.2 years	2.75 years
Average Annual Production	1,478 tonnes TREO (ex La, Ce), including: 119 t Dy, 17.8 t Tb, 841 t Y	Same as Base Case
Mine Plan	32 million tonnes Proven and Probable Reserves	Same as Base Case
Estimated Life of Mine	13-year mine life	Same as Base Case

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Metric	Base Case	Divergent Pricing
Rare Earth Oxide Prices Used (average Life of Mine)	Dy₂O₃: USD663/kg
Tb₂O₃: USD2,880/kg
Y₂O₃: USD60/kg
Nd₂O₃: USD114/kg
Pr₆O₁₁: USD119/kg	Dy₂O₃: USD855/kg
Tb₂O₃: USD3,712/kg
Y₂O₃: USD130/kg
Nd₂O₃: USD146/kg
Pr₆O₁₁: USD152/kg
Basket Price (average Life of Mine pricing)	USD158/kg excluding La,Ce	USD230/kg excluding La,Ce

22.2. Basis of Economic Analysis

The analysis has been conducted on a pre-debt financing basis. Escalation and inflation have been excluded.

The currency adopted for the analysis is the USD.

In calculating the returns from the project, the following fundamental assumptions have been made:

The operating life of the project will be 13 years;
The design throughput for the process plant is 1 482 000 tonnes per year,
The economic returns are assessed at the project level on a pre-financing basis;
The evaluation includes a 24-month project development period prior to the commencement of production, giving the project a total life of 15 years;
The exchange rate used to convert ZAR into USD is 18.23 to 1. The exchange rate used to convert NAD into USD is 18.23 to 1.

All assumptions made as part of the economic evaluation are detailed in Section 22.2.3.

22.2.1. Project Contingency

Contingency is applied to the capital costs of the project, based on the level of engineering design, construction factors and vendor quotes used for the capital cost estimate.

Contingency is applied as shown in Table 145.

Table 145 Project Contingency

Area	Amount	Factor
Process Plant	$41 478 323	32% of Total Installed Costs
Non Process Infrastructure	$11 749 316	20% of Direct Field Costs
Tailings	$4 132 428	18% of Installation, Operation and Closure Costs
Total Contingency	$57 360 067	19.7% of Total Capital Costs

Mining capital shows no contingency, as the contractor quotes have several different factors built into their pricing and thus couldn't be readily derived from their costs.

Overall, a 20% contingency for the project is deemed reasonable at the PFS stage.

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22.2.2. Rare Earth Pricing Models

Two pricing models are presented in the PFS:

Divergent Case - consistent with consensus divergence pricing logic and the geopolitical reality of prolonged export controls, slow non-Chinese separation build-out and rising strategic demand from OEMs.
Base Case – representing a conservative but realistic post-normalization scenario, assuming partial easing of China's recent export bottlenecks.

Table 146 Rare Earth Oxide Pricing

Rare Earth Oxide Prices Used (average Life of Mine)	Base Case [USD/kg]	Divergent Case [USD/kg]
Dy_{203}	663	855
Tb_{203}	2,880	3,712
Y_{203}	60	130
Nd_{203}	114	146
Pr_{6011}	119	152
Basket Price (average Life of Mine pricing)	USD158/kg excluding LaCe	USD230/kg excluding LaCe

22.2.3. Clarification and Assumptions

Economic analysis has been carried out on the basis detailed within this section.

22.2.4. Analysis Period

The period of analysis is from the commencement of Process plant construction during quarter one, Year -2 until the end of mining through to the last sale of product in Year 13.

22.2.5. Operating Costs

Operating costs have been estimated as per the following functional headings:

Mining;
Mill Site Process Plant;
General and Administration.

A full description of operating costs can be found in Section 21 of the Report.

22.2.6. Capital Costs

Capital costs include the direct capital costs the process plant, non process infrastructure and tailings storage facility; sustaining capital for the process plant; mine and tailings facility closure

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costs; indirect costs (including EPCM, first fills, spares and a construction camp allowance) and a contingency.

A full description of capital costs can be found in Section 21.1 of the Report.

22.2.7. Funding

The financial model for the project is presented on the basis of pre-financing cash flows and as such excludes the impact of both debt funding and equity finance.

22.2.8. After Tax Free Cash Flow

After-tax free cash flow is calculated by deducting operating costs, royalties, taxes and sustaining capital expenditures from revenue.

22.2.9. Net Present Value

After-tax NPVs are calculated from the annual free cash flows. The financial model can apply a range of discount factors. The use of various discount rates in the base case and sensitivity analysis of this report should not be taken as an endorsement of those discount rates as appropriate rates of return for this project.

22.2.10. After Tax Internal Rate of Return

The after-tax IRR is calculated from the annual after-tax free cash flows.

22.2.11. Payback Period

The payback period is identified as the period in which the cumulative undiscounted (after-tax free) cash flow becomes positive, having paid back the development costs.

22.3. Base Case Financial Model and Sensitivities

22.3.1. Base Case Project Economics

Based on the extraction of 3 010 000 t/a of ROM feed from the mine and process feed of 1 482 000 tonnes per year, the project is anticipated to yield a pre-tax IRR of 21.7% with a pre-tax NPV, at a discount rate of 5% of USD $389.2M, and an after-tax IRR of 19% with an after-tax NPV, at a discount rate of 5%, of USD $275.5M. Cumulative cash flows are USD $710M M pre-tax and USD $513.1 M after-tax over the thirteen year LOM.

The project is expected to pay back initial capital in 4.2 years after production starts.

Figure 203 shows the cumulative after-tax cash flow over the total project life.

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Figure 203 Base Case After Tax Cash Flows

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Table 147 Base Case Financial Model

		-2	-1	1	2	3	4	5	6	7	8	9	10	11	12	13
Final Recovered Metals
Le2O3	kg	2,583k		58k	149k	196k	218k	223k	191k	269k	196k	253k	255k	189k	198k	187k
Ce2O3	kg	4,528k		105k	265k	345k	381k	395k	347k	445k	340k	443k	438k	334k	366k	325k
Pr2O3	kg	477k		11k	28k	36k	40k	42k	37k	45k	36k	46k	46k	35k	40k	34k
Ne2O3	kg	1,731k		42k	105k	130k	144k	151k	138k	161k	132k	167k	169k	126k	148k	118k
Sm2O3	kg	559k		17k	36k	44k	45k	47k	52k	46k	47k	53k	58k	36k	46k	32k
Eu2O3	kg	253k		9k	17k	22k	22k	21k	27k	19k	22k	24k	24k	15k	19k	12k
Gd2O3	kg	1,063k		43k	76k	100k	95k	89k	119k	71k	91k	100k	95k	58k	80k	44k
Th2O3	kg	232k		11k	17k	24k	22k	19k	28k	14k	19k	22k	19k	11k	17k	8k
Dy2O3	kg	1,553k		77k	118k	170k	149k	129k	196k	84k	126k	146k	121k	66k	119k	53k
Ho2O3	kg	329k		17k	25k	37k	32k	27k	43k	17k	26k	31k	24k	13k	25k	11k
Er2O3	kg	959k		52k	73k	110k	95k	79k	127k	46k	74k	90k	69k	37k	74k	31k
Tm2O3	kg	142k		8k	11k	17k	14k	12k	19k	7k	11k	13k	10k	5k	11k	4k
Yb2O3	kg	861k		49k	65k	101k	85k	72k	117k	40k	65k	80k	60k	31k	68k	27k
Lu2O3	kg	124k		7k	9k	15k	12k	10k	17k	6k	9k	12k	9k	4k	10k	4k
Y2O3	kg	10,937k		581k	827k	1,252k	1,113k	916k	1,481k	523k	838k	1,053k	794k	414k	809k	337k
Total LREO Recovered		9,879k		234k	583k	752k	827k	858k	765k	966k	750k	963k	966k	720k	797k	697k
Total HREO Recovered		16,453k		856k	1,240k	1,849k	1,639k	1,375k	2,173k	826k	1,281k	1,571k	1,225k	654k	1,233k	531k
Total TREO Recovered		26,332k		1,090k	1,823k	2,601k	2,466k	2,232k	2,939k	1,792k	2,032k	2,534k	2,192k	1,374k	2,030k	1,228k
Operating Costs
Mining		$652.4 M		$32.8 M	$45.2 M	$52.8 M	$57.6 M	$57.3 M	$56.7 M	$59.7 M	$60.9 M	$69.0 M	$55.6 M	$48.6 M	$39.5 M	$16.8 M
Processing		$996.3 M		$40.0 M	$71.0 M	$82.7 M	$84.4 M	$84.4 M	$84.5 M	$84.4 M	$84.4 M	$84.4 M	$84.4 M	$67.8 M	$84.4 M	$59.3 M
G&A		$29.8 M		$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M	$2.3 M
Total Costs		$1,678.5 M		$75.1 M	$118.5 M	$137.7 M	$144.3 M	$144.0 M	$143.4 M	$146.4 M	$147.7 M	$155.7 M	$142.4 M	$118.8 M	$126.3 M	$78.4 M
Revenues / Royalties
Overall Gross Revenue		$3,031.4 M		$122.9 M	$198.2 M	$295.9 M	$275.5 M	$249.7 M	$381.9 M	$183.0 M	$256.6 M	$304.1 M	$254.4 M	$144.6 M	$246.2 M	$118.5 M
Royalties and Separation Costs		$299.4 M		$13.1 M	$20.4 M	$30.3 M	$27.8 M	$24.6 M	$37.1 M	$17.3 M	$24.0 M	$29.0 M	$23.9 M	$13.6 M	$23.2 M	$11.3 M
Gross Revenue Net of Royalties		$2,736.0 M		$109.8 M	$177.8 M	$265.6 M	$247.7 M	$225.1 M	$344.8 M	$165.7 M	$232.6 M	$275.1 M	$230.5 M	$131.0 M	$223.0 M	$107.2 M
Net Revenue After Costs		$1,057.5 M		$0.0 M	$34.8 M	$59.4 M	$127.9 M	$103.4 M	$81.1 M	$201.4 M	$19.3 M	$85.0 M	$119.4 M	$88.1 M	$12.2 M	$96.8 M
TREO Realized Revenue Per KG (Basket)					$38	$43	$63	$57	$52	$86	$19	$59	$67	$61	$15	$68
Capital Costs
Mining Capital		$27.6 M	$0.0 M	$21.3 M	$0.6 M	$3.3 M	$0.1 M	$0.0 M	$0.4 M	$0.1 M	$0.0 M	$0.3 M	$0.0 M	$0.3 M	-$0.1 M	$0.1 M
Process capital		$181.6 M	$74.8 M	$92.7 M	$7.5 M	$6.6 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M
Facilities Capital		$58.7 M	$28.0 M	$29.6 M	$1.1 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M
Tailings Capital		$21.6 M	$0.0 M	$10.9 M	$0.2 M	$8.5 M	$0.2 M	$0.2 M	$0.2 M	$0.2 M	$0.2 M	$0.2 M	$0.2 M	$0.2 M	$0.4 M	$0.0 M
Closure Costs		$1.0 M	$0.0 M	$0.0 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.2 M
Total Capital		$290.6 M	$102.9 M	$154.4 M	$9.5 M	$18.5 M	$0.4 M	$0.3 M	$0.6 M	$0.3 M	$0.3 M	$0.5 M	$0.3 M	$0.5 M	$0.2 M	$0.4 M
Total Project Contingency		$57.4 M	$0.0 M	$0.0 M	$54.9 M	$0.1 M	$1.3 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.1 M	$0.2 M
Net Cash Flow		$709.6 M	-$102.9 M	-$154.4 M	-$29.6 M	$40.8 M	$126.2 M	$103.0 M	$80.4 M	$201.0 M	$19.0 M	$84.3 M	$119.0 M	$87.5 M	$12.0 M	$96.3 M
Cumulative Net Cash Flow			-$102.9 M	-$257.3 M	-$286.9 M	-$246.1 M	-$119.9 M	-$16.9 M	$63.5 M	$264.5 M	$283.4 M	$367.8 M	$486.8 M	$574.3 M	$586.2 M	$682.5 M
Pre-Tax NPV			$389.2 M

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		-2	-1	1	2	3	4	5	6	7	8	9	10	11	12	13
Pre-Tax IRR		21.7%
Operating Income		$0.0 M	$0.0 M	$34.8 M	$59.4 M	$127.9 M	$103.4 M	$81.1 M	$201.4 M	$19.3 M	$85.0 M	$119.4 M	$88.1 M	$12.2 M	$96.8 M	$28.8 M
Development Capex		-$102.9 M	-$154.4 M	-$9.5 M	-$18.5 M	-$0.4 M	-$0.3 M	-$0.6 M	-$0.3 M	-$0.3 M	-$0.5 M	-$0.3 M	-$0.3 M	-$0.2 M	-$0.4 M	-$1.6 M
Taxable Income and capital costs current year		-$102.9 M	-$154.4 M	$25.3 M	$40.9 M	$127.5 M	$103.1 M	$80.5 M	$201.1 M	$19.1 M	$84.4 M	$119.1 M	$87.6 M	$12.1 M	$96.4 M	$27.2 M
Carry forward		-$24.5 M	-$127.3 M	-$367.6 M	-$421.7 M	-$423.2 M	-$304.8 M	-$207.7 M	-$127.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M
Taxable income after carry forward		-$127.3 M	-$281.8 M	-$332.8 M	-$362.4 M	-$295.3 M	-$201.4 M	-$126.6 M	$74.4 M	$19.3 M	$85.0 M	$119.4 M	$88.1 M	$12.2 M	$96.8 M	$28.8 M
Carry forward of exploration expense		-$24.5 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M
Capital Allowance for development expense (3 years)		-$34.3 M	-$85.8 M	-$88.9 M	-$60.8 M	-$9.4 M	-$6.4 M	-$0.4 M	-$0.4 M	-$0.4 M	-$0.4 M	-$0.3 M	-$0.4 M	-$0.3 M	-$0.4 M	-$0.7 M
Taxable income after carry forward and capital allowance		-$58.8 M	-$367.6 M	-$421.7 M	-$423.2 M	-$304.8 M	-$207.7 M	-$127.0 M	$73.9 M	$18.9 M	$84.6 M	$119.1 M	$87.7 M	$11.9 M	$96.4 M	$28.1 M
Tax Rate			$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M
Net Tax Payable		$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$27.9 M	$7.2 M	$31.9 M	$44.8 M	$33.0 M	$4.6 M	$36.3 M	$10.8 M
After Tax Cash Flow	$513.1 M	-$102.9 M	-$154.4 M	-$29.6 M	$40.8 M	$126.2 M	$103.0 M	$80.4 M	$173.1 M	$11.7 M	$52.5 M	$74.2 M	$54.4 M	$7.4 M	$60.0 M	$16.3 M
After Tax Cumulative Cash Flow		-$102.9 M	-$257.3 M	-$286.9 M	-$246.1 M	-$119.9 M	-$16.9 M	$63.5 M	$236.6 M	$248.3 M	$300.8 M	$375.0 M	$429.5 M	$436.8 M	$496.8 M	$513.1 M
After Tax NPV		$275.5 M
After Tax IRR		19.0%

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22.3.2. Base Case Sensitivity Analysis

For the purposes of this PFS, the evaluation is based on 100% of the Project cash flows before distribution of profits to equity owners. Economic sensitivities are presented for various scenarios:

All Values are After Tax
Discount rates of 5%, 6, 7%, 8%, 9%, 10% and 11%
Sensitivity ranges for operating and capital costs between +/- 20% of base case values
Sensitivity ranges for TREO recoveries from 48% to 72%
Sensitivity ranges for revenues (Basket Pricing) of -20% to +20%

Table 148 to Table 152 present the various sensitivity results. Figure 204 shows the Sensitivity Graph for the various ranges.

Table 148 Capital Cost Sensitivity

Discount Rate	80%	90%	100%	110%	120%
5%	$312.4 M	$294.0 M	$275.5 M	$257.1 M	$238.2 M
6%	$278.7 M	$259.9 M	$241.1 M	$222.4 M	$203.1 M
7%	$248.2 M	$229.1 M	$210.1 M	$191.0 M	$171.4 M
8%	$220.6 M	$201.3 M	$182.0 M	$162.7 M	$142.8 M
9%	$195.5 M	$176.0 M	$156.5 M	$137.0 M	$117.0 M
10%	$172.6 M	$153.0 M	$133.4 M	$113.8 M	$93.6 M
11%	$151.9 M	$132.2 M	$112.4 M	$92.7 M	$72.4 M

Table 149 Operating Cost Sensitivity

Discount Rate	80%	90%	100%	110%	120%
5%	$417.2 M	$346.4 M	$275.5 M	$204.6 M	$131.4 M
6%	$373.0 M	$307.0 M	$241.1 M	$175.2 M	$106.7 M
7%	$332.9 M	$271.5 M	$210.1 M	$148.6 M	$84.4 M
8%	$296.7 M	$239.3 M	$182.0 M	$124.6 M	$64.3 M
9%	$263.8 M	$210.1 M	$156.5 M	$102.8 M	$46.2 M
10%	$233.9 M	$183.6 M	$133.4 M	$83.1 M	$29.8 M
11%	$206.7 M	$159.6 M	$112.4 M	$65.2 M	$15.0 M

Table 150 Exchange Rate Sensitivity

NAD:USD Exchange	NAD 14.58	NAD 16.40	NAD 18.23	NAD 20.05	NAD 21.87
Discount Rate	80%	90%	100%	110%	120%
5%	$125.6 M	$210.4 M	$275.5 M	$328.5 M	$372.3 M
6%	$97.9 M	$179.2 M	$241.1 M	$291.5 M	$333.2 M
7%	$73.1 M	$151.0 M	$210.1 M	$258.1 M	$297.8 M
8%	$50.7 M	$125.5 M	$182.0 M	$227.8 M	$265.8 M
9%	$30.6 M	$102.4 M	$156.5 M	$200.3 M	$236.6 M
10%	$12.5 M	$81.6 M	$133.4 M	$175.4 M	$210.2 M
11%	-$3.9 M	$62.7 M	$112.4 M	$152.7 M	$186.1 M

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Table 151 Base Case Metal Price Sensitivity

Discount Rate	80%	90%	100%	110%	120%
5%	-$10.0 M	$144.6 M	$275.5 M	$404.4 M	$506.1 M
6%	-$25.7 M	$118.9 M	$241.1 M	$361.2 M	$455.8 M
7%	-$39.7 M	$95.7 M	$210.1 M	$322.1 M	$410.3 M
8%	-$52.1 M	$74.8 M	$182.0 M	$286.6 M	$369.0 M
9%	-$63.2 M	$55.9 M	$156.5 M	$254.5 M	$331.6 M
10%	-$73.1 M	$38.9 M	$133.4 M	$225.3 M	$297.5 M
11%	-$81.8 M	$23.5 M	$112.4 M	$198.7 M	$266.6 M

Table 152 Recovery Sensitivity

LREO recovery	48.0%	50.9%	53.7%	56.5%	59.3%	62.2%	65.0%
HREO Recovery	52.9%	56.0%	59.1%	62.2%	65.4%	68.5%	71.6%
Discount Rate	85%	90%	95%	100%	105%	110%	115%
5%	$97.0 M	$157.9 M	$217.4 M	$275.5 M	$333.6 M	$391.7 M	$442.5 M
6%	$74.2 M	$131.4 M	$187.0 M	$241.1 M	$295.2 M	$349.3 M	$396.6 M
7%	$53.7 M	$107.5 M	$159.6 M	$210.1 M	$260.5 M	$311.0 M	$355.1 M
8%	$35.3 M	$85.9 M	$134.8 M	$182.0 M	$229.1 M	$276.3 M	$317.5 M
9%	$18.8 M	$66.4 M	$112.4 M	$156.5 M	$200.6 M	$244.8 M	$283.3 M
10%	$3.9 M	$48.8 M	$92.0 M	$133.4 M	$174.8 M	$216.1 M	$252.3 M
11%	-$9.5 M	$32.8 M	$73.6 M	$112.4 M	$151.3 M	$190.1 M	$224.0 M

Figure 204 Base Case After Tax Sensitivity

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22.4. Divergent Pricing Financial Model and Sensitivities

22.4.1. Divergent Pricing Project Economics

Based on the extraction of 3 010 000 t/a of ROM feed from the mine and process feed of 1 482 000 tonnes per year, the project is anticipated to yield a pre-tax IRR of 44.1% with a pre-tax NPV, at a discount rate of 5% of USD $1 245.6M, and an after-tax IRR of 34.8% with an after-tax NPV, at a discount rate of 5%, of USD $747.9M. Cumulative cash flows are USD $2 027M M pre-tax and USD $1 242.3 M after-tax over the thirteen year LOM.

The project is expected to pay back initial capital in 2.75 years after production starts.

Figure 205 shows the cumulative after-tax cash flow over the total project life.

Figure 205 Divergent Pricing After Tax Cash Flows

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Table 153 Divergent Pricing Financial Model

				-2	-1	1	2	3	4	5	6	7	8	9	10	11	12	13
Final Recovered Metals
	Le2O3	kg	2,583k			58k	149k	196k	218k	223k	191k	269k	196k	253k	255k	189k	198k	187k
	Ce2O3	kg	4,528k			105k	265k	345k	381k	395k	347k	445k	340k	443k	438k	334k	366k	325k
	Pr2O3	kg	477k			11k	28k	36k	40k	42k	37k	45k	36k	46k	46k	35k	40k	34k
	Nd2O3	kg	1,731k			42k	105k	130k	144k	151k	138k	161k	132k	167k	169k	126k	148k	118k
	Sm2O3	kg	559k			17k	36k	44k	45k	47k	52k	46k	47k	53k	58k	36k	46k	32k
	Eu2O3	kg	253k			9k	17k	22k	22k	21k	27k	19k	22k	24k	24k	15k	19k	12k
	Gd2O3	kg	1,063k			43k	76k	100k	95k	89k	119k	71k	91k	100k	95k	58k	80k	44k
	Tb2O3	kg	232k			11k	17k	24k	22k	19k	28k	14k	19k	22k	19k	11k	17k	8k
	Dy2O3	kg	1,553k			77k	118k	170k	149k	129k	196k	84k	126k	146k	121k	66k	119k	53k
	Ho2O3	kg	329k			17k	25k	37k	32k	27k	43k	17k	26k	31k	24k	13k	25k	11k
	Er2O3	kg	959k			52k	73k	110k	95k	79k	127k	46k	74k	90k	69k	37k	74k	31k
	Tm2O3	kg	142k			8k	11k	17k	14k	12k	19k	7k	11k	13k	10k	5k	11k	4k
	Yb2O3	kg	861k			49k	65k	101k	85k	72k	117k	40k	65k	80k	60k	31k	68k	27k
	Lu2O3	kg	124k			7k	9k	15k	12k	10k	17k	6k	9k	12k	9k	4k	10k	4k
	Y2O3	kg	10,937k			581k	827k	1,252k	1,113k	916k	1,481k	523k	838k	1,053k	794k	414k	809k	337k
	Total UREO Recovered		9,879k			234k	583k	752k	827k	858k	765k	966k	750k	963k	966k	720k	797k	697k
	Total HREO Recovered		16,453k			856k	1,240k	1,849k	1,639k	1,375k	2,173k	826k	1,281k	1,571k	1,225k	654k	1,233k	531k
	Total TREO Recovered		26,332k			1,090k	1,823k	2,601k	2,466k	2,232k	2,939k	1,792k	2,032k	2,534k	2,192k	1,374k	2,030k	1,228k
Operating Costs
	Mining		5652.4 M			532.8 M	545.2 M	552.8 M	557.6 M	557.3 M	556.7 M	559.7 M	560.9 M	569.0 M	555.6 M	548.6 M	539.5 M	516.8 M
	Processing		5996.3 M			540.0 M	571.0 M	582.7 M	584.4 M	584.4 M	584.5 M	584.4 M	584.4 M	584.4 M	584.4 M	567.8 M	584.4 M	559.3 M
	G&A		529.8 M			52.3 M	52.3 M	52.3 M	52.3 M	52.3 M	52.3 M	52.3 M	52.3 M	52.3 M	52.3 M	52.3 M	52.3 M	52.3 M
	Total Costs		51,678.5 M			575.1 M	5118.5 M	5137.7 M	5144.3 M	5144.0 M	5143.4 M	5146.4 M	5147.7 M	5155.7 M	5142.4 M	5118.8 M	5126.3 M	578.4 M
Revenues / Royalties
	Overall Gross Revenue		54,418.6 M			5167.4 M	5271.5 M	5415.8 M	5394.0 M	5360.8 M	5569.7 M	5268.6 M	5382.9 M	5457.3 M	5377.1 M	5212.1 M	5367.7 M	5173.8 M
	Royalties and Separation Costs		5364.7 M			515.3 M	524.1 M	536.3 M	533.7 M	530.1 M	546.5 M	521.5 M	530.3 M	536.6 M	530.0 M	516.9 M	529.3 M	514.0 M
	Gross Revenue Net of Royalties		54,053.9 M			5152.1 M	5247.4 M	5379.6 M	5360.2 M	5330.7 M	5523.2 M	5247.1 M	5352.6 M	5420.7 M	5347.1 M	5195.1 M	5338.4 M	5159.7 M
	Net Revenue After Costs		52,375.3 M			577.0 M	5128.9 M	5241.8 M	5215.9 M	5186.6 M	5379.8 M	5100.7 M	5204.9 M	5265.0 M	5204.8 M	576.3 M	5212.2 M	581.3 M
	TREO Realized Revenue Per KG (Basket)					584.67	593.85	5120.00	5118.47	5119.11	5161.75	597.57	5141.44	5148.47	5142.11	595.71	5149.51	5119.17
Capital Costs
	Mining Capital		527.6 M	50.0 M	521.3 M	50.6 M	53.3 M	50.1 M	50.0 M	50.4 M	50.1 M	50.0 M	50.3 M	50.0 M	50.3 M	-50.1 M	-50.1 M	51.4 M
	Process capital		5181.6 M	574.8 M	592.7 M	57.5 M	56.6 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M
	Facilities Capital		558.7 M	528.0 M	529.6 M	51.1 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M	50.0 M
	Tailings Capital		521.6 M	50.0 M	510.9 M	50.2 M	58.5 M	50.2 M	50.2 M	50.2 M	50.2 M	50.2 M	50.2 M	50.2 M	50.2 M	50.2 M	50.2 M	50.2 M
	Closure Costs		51.0 M	50.0 M	50.0 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.2 M
	Total Capital		5290.6 M	5102.9 M	5154.4 M	59.5 M	518.5 M	50.4 M	50.3 M	50.6 M	50.3 M	50.3 M	50.5 M	50.3 M	50.5 M	50.2 M	50.4 M	51.6 M
	Total Project Contingency		557.4 M	50.0 M	50.0 M	554.9 M	50.1 M	51.3 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.1 M	50.2 M
	Net Cash Flow		52,027.4 M	-5102.9 M	-5154.4 M	512.7 M	5110.3 M	5240.2 M	5215.5 M	5185.9 M	5379.4 M	5100.3 M	5204.3 M	5264.6 M	5204.1 M	576.1 M	5211.7 M	579.6 M
	Cumulative Net Cash Flow			-5102.9 M	-5257.3 M	-5244.6 M	-5134.3 M	5105.9 M	5321.4 M	5507.3 M	5886.7 M	5987.1 M	51,191.3 M	51,455.9 M	51,660.1 M	51,736.1 M	51,947.8 M	52,027.4 M
	Pre-Tax NPV			51,245.6 M

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	-2	-1	1	2	3	4	5	6	7	8	9	10	11	12	13
Pre-Tax IRR	44.1%
Operating Income	$0.0 M	$0.0 M	$77.0 M	$128.9 M	$241.8 M	$215.9 M	$186.6 M	$379.8 M	$100.7 M	$204.9 M	$265.0 M	$204.8 M	$76.3 M	$212.2 M	$81.3 M
Development Capex	-$102.9 M	-$154.4 M	-$9.5 M	-$18.5 M	-$0.4 M	-$0.3 M	-$0.6 M	-$0.3 M	-$0.3 M	-$0.5 M	-$0.3 M	-$0.5 M	-$0.2 M	-$0.4 M	-$1.6 M
Taxable Income and capital costs current year	-$102.9 M	-$154.4 M	$67.5 M	$110.4 M	$241.5 M	$215.6 M	$186.0 M	$379.5 M	$100.4 M	$204.4 M	$264.7 M	$204.2 M	$76.2 M	$211.8 M	$79.8 M
Carry forward	-$24.5 M	-$127.3 M	-$281.8 M	-$204.8 M	-$75.9 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M
Taxable income after carry forward	-$127.3 M	-$281.8 M	-$204.8 M	-$75.9 M	$165.9 M	$215.9 M	$186.6 M	$379.8 M	$100.7 M	$204.9 M	$265.0 M	$204.8 M	$76.3 M	$212.2 M	$81.3 M
Carry forward of exploration expense	-$24.5 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M
Capital Allowance for development expense (3 years)	-$34.3 M	-$85.8 M	-$88.9 M	-$60.8 M	-$9.4 M	-$6.4 M	-$0.4 M	-$0.4 M	-$0.4 M	-$0.4 M	-$0.3 M	-$0.4 M	-$0.3 M	-$0.4 M	-$0.7 M
Taxable income after carry forward and capital allowance	-$58.8 M	-$85.8 M	-$293.7 M	-$136.7 M	$156.5 M	$209.6 M	$186.2 M	$379.4 M	$100.3 M	$204.5 M	$264.6 M	$204.3 M	$76.0 M	$211.8 M	$80.6 M
Tax Rate		$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$0.0 M
Net Tax Payable	$0.0 M	$0.0 M	$0.0 M	$0.0 M	$62.2 M	$81.0 M	$70.0 M	$142.4 M	$37.8 M	$76.8 M	$99.4 M	$76.8 M	$28.6 M	$79.6 M	$30.5 M
After Tax Cash Flow	$1,242.3 M	-$102.9 M	-$154.4 M	$12.7 M	$110.3 M	$178.0 M	$134.6 M	$115.9 M	$237.0 M	$62.6 M	$127.4 M	$165.2 M	$127.3 M	$47.4 M	$132.1 M
After Tax Cumulative Cash Flow		-$102.9 M	-$257.3 M	-$244.6 M	-$134.3 M	$43.7 M	$178.2 M	$294.1 M	$531.1 M	$593.7 M	$721.1 M	$886.3 M	$1,013.7 M	$1,061.1 M	$1,193.3 M
After Tax NPV		$747.9 M
After Tax IRR		34.8%

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22.4.2. Divergent Pricing Sensitivity Analysis

For the purposes of this PFS, the evaluation is based on 100% of the Project cash flows before distribution of profits to equity owners. Economic sensitivities are presented for various scenarios:

All Values are After Tax
Discount rates of 5%, 6, 7%, 8%, 9%, 10% and 11%
Sensitivity ranges for operating and capital costs between +/- 20% of base case values
Sensitivity ranges for TREO recoveries from 48% to 72%
Sensitivity ranges for revenues (Basket Pricing) of -20% to +20%

Table 154 to Table 158 present the various sensitivity results. Figure 206 shows the Sensitivity Graph for the various ranges.

Table 154 Capital Cost Sensitivity

Discount Rate	80%	90%	100%	110%	120%
5%	$795.5 M	$771.7 M	$747.9 M	$724.1 M	$700.2 M
6%	$723.9 M	$700.4 M	$676.8 M	$653.3 M	$629.7 M
7%	$659.3 M	$636.0 M	$612.7 M	$589.4 M	$566.1 M
8%	$600.8 M	$577.8 M	$554.7 M	$531.7 M	$508.7 M
9%	$547.8 M	$525.0 M	$502.3 M	$479.5 M	$456.7 M
10%	$499.7 M	$477.2 M	$454.7 M	$432.1 M	$409.6 M
11%	$456.0 M	$433.7 M	$411.4 M	$389.1 M	$366.8 M

Table 155 Operating Cost Sensitivity

Discount Rate	80%	90%	100%	110%	120%
5%	$795.5 M	$771.7 M	$747.9 M	$724.1 M	$700.2 M
6%	$723.9 M	$700.4 M	$676.8 M	$653.3 M	$629.7 M
7%	$659.3 M	$636.0 M	$612.7 M	$589.4 M	$566.1 M
8%	$600.8 M	$577.8 M	$554.7 M	$531.7 M	$508.7 M
9%	$547.8 M	$525.0 M	$502.3 M	$479.5 M	$456.7 M
10%	$499.7 M	$477.2 M	$454.7 M	$432.1 M	$409.6 M
11%	$456.0 M	$433.7 M	$411.4 M	$389.1 M	$366.8 M

Table 156 Exchange Rate Sensitivity

NAD:USD Exchange	NAD 14.58	NAD 16.40	NAD 18.23	NAD 20.05	NAD 21.87
Discount Rate	80%	90%	100%	110%	120%
5%	$795.5 M	$771.7 M	$747.9 M	$724.1 M	$700.2 M
6%	$723.9 M	$700.4 M	$676.8 M	$653.3 M	$629.7 M
7%	$659.3 M	$636.0 M	$612.7 M	$589.4 M	$566.1 M
8%	$600.8 M	$577.8 M	$554.7 M	$531.7 M	$508.7 M
9%	$547.8 M	$525.0 M	$502.3 M	$479.5 M	$456.7 M
10%	$499.7 M	$477.2 M	$454.7 M	$432.1 M	$409.6 M
11%	$456.0 M	$433.7 M	$411.4 M	$389.1 M	$366.8 M

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Table 157 Divergent Pricing Metal Price Sensitivity

Discount Rate	80%	90%	100%	110%	120%
5%	$795.5 M	$771.7 M	$747.9 M	$724.1 M	$700.2 M
6%	$723.9 M	$700.4 M	$676.8 M	$653.3 M	$629.7 M
7%	$659.3 M	$636.0 M	$612.7 M	$589.4 M	$566.1 M
8%	$600.8 M	$577.8 M	$554.7 M	$531.7 M	$508.7 M
9%	$547.8 M	$525.0 M	$502.3 M	$479.5 M	$456.7 M
10%	$499.7 M	$477.2 M	$454.7 M	$432.1 M	$409.6 M
11%	$456.0 M	$433.7 M	$411.4 M	$389.1 M	$366.8 M

Table 158 Process Recovery Sensitivity

LREO recovery	48.0%	50.9%	53.7%	56.5%	59.3%	62.2%	65.0%
HREO Recovery	52.9%	56.0%	59.1%	62.2%	65.4%	68.5%	71.6%
Discount Rate	85%	90%	95%	100%	105%	110%	115%
5%	$497.7 M	$581.1 M	$664.5 M	$747.9 M	$831.3 M	$914.7 M	$998.1 M
6%	$445.4 M	$522.5 M	$599.7 M	$676.8 M	$754.0 M	$831.1 M	$908.3 M
7%	$398.2 M	$469.7 M	$541.2 M	$612.7 M	$684.2 M	$755.7 M	$827.2 M
8%	$355.6 M	$422.0 M	$488.4 M	$554.7 M	$621.1 M	$687.5 M	$753.9 M
9%	$317.1 M	$378.8 M	$440.5 M	$502.3 M	$564.0 M	$625.7 M	$687.4 M
10%	$282.2 M	$339.7 M	$397.2 M	$454.7 M	$512.1 M	$569.6 M	$627.1 M
11%	$250.6 M	$304.2 M	$357.8 M	$411.4 M	$465.0 M	$518.7 M	$572.3 M

Figure 206 Divergent Pricing Sensitivities

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22.4.3. Recovery Sensitivities

The sensitivity graph for process recoveries is shown in Figure 22-5. This graph shows the recovery sensitivity for both the base case pricing and the divergent case pricing.

Figure 207 Process Recovery Sensitivities

22.4.4. Discussion on Sensitivities

The sensitivity tables and charts show that the project is reasonably robust with the only negative values occurring at the -20% range for metal prices under the base case scenario, and at the -20% range and 11% discount rate for USD exchange rates and process recoveries.

Under Divergent Pricing, all sensitivity ranges are positive.

The project is most sensitive to metal prices, followed by Process Recoveries, then by USD exchange rates. The project shows lower sensitivities to operating and capital costs.

22.4.5. Conclusions and Recommendations

The project shows robust economics with positive net present values across both pricing cases and sensitivity ranges, only exhibiting negative values at the lowest range of recoveries, metal prices, and high NAD:USD exchanges, on the base case rare earth oxide pricing.

Project economics may be improved by developing a full geo-metallurgy model to relate reagent consumption costs against grades, potentially Improving project economics

It is recommended that the project proceed to full feasibility study.

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23. Adjacent Properties

There are no adjacent properties of relevance to this report.

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24. Other Relevant Data and Information

There is no other relevant data or information available that is necessary to make the current technical report understandable and not misleading.

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25. Interpretation and Conclusions

25.1. Mineral Resource Estimate

On behalf of NMI, MSA has completed an update to the Mineral Resource estimates for Area 4 and Area 2B of the Lofdal Heavy Rare Earths Project.

The Mineral Resource is reported as Measured, Indicated and Inferred Mineral Resources as shown in Table 159 for Area 4 and Table 160 for Area 2B. The Mineral Resource was estimated using The Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Best Practice Guidelines (2019) and is reported in accordance with the 2014 CIM Definition Standards, which have been incorporated by reference into National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101).

In the QP’s opinion, the Mineral Resources reported herein at the selected cut-off grade have “reasonable prospects for eventual economic extraction”, taking into consideration mining and processing assumptions (refer to Table 159). The Mineral Resource was reported from within a Whittle optimised pit shell at a cut-off grade of 0.10% TREO.

Table 159 Area 4 Mineral Resources above a 0.1% TREO cut-off grade – 5 April 2024

Category	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy₂O₃ ppm	TREO (kt)
Measured	6.6	0.21	0.14	0.07	130	13.7
Indicated	49.2	0.15	0.07	0.08	69	75.7
Measured & Indicated	55.8	0.16	0.08	0.08	76	89.4
Inferred	10.5	0.14	0.06	0.08	58	15.0

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Table 160 Area 2B Mineral Resources above a 0.1% TREO cut-off grade – 5 April 2024

Category	Tonnes (Mt)	TREO* %	HREO** %	LREO*** %	Dy₂O₃ ppm	TREO (kt)
Indicated	2.7	0.16	0.09	0.07	97	4.4
Inferred	4.4	0.15	0.07	0.08	75	6.6

Notes:
8. All tabulated data have been rounded and as a result minor computational errors may occur.
9. Mineral Resources, which are not Mineral Reserves, have no demonstrated economic viability.
10. Quantities reported are the total quantities for the project regardless of ownership.
11. TREO = Total Rare Earth Oxides and includes Y2O3
12. HREO = Heavy Rare Earth Oxides and includes Y2O3
13. **LREO = Light Rare Earth Oxides
14. Mt = Million tonnes, kt = Thousand tonnes,.

The Area 4 Mineral Resource Estimate has increased from the previous estimate of 12 May 2021, largely due changes to the input parameters used to generate the optimised shell which differ to those used in the previous Mineral Resource estimate. The Area 2B Indicated Mineral Resource has seen a slight increase in tonnages due to changes to the input parameters for the optimised pit shell. The Inferred Mineral Resources have increased in tonnages due to the additional RC drilling which has extended the Mineral Resource in a northeasterly direction along strike.

25.2. Capital and Operating Cost

Process capital includes the process plant and ore sorting facility.

Facilities capital includes all non-process site facilities, including water and power supply, non-process site buildings, security and warehousing.

CAPEX increases reflect inflation since the 2022 PEA, expanded hydrometallurgical scope (acid recovery), inclusion of mining pre-strip and revised power infrastructure requirements.

Table 161 Capital Costs Summary of Lofdal PFS "Lofdal 2B-4"

Capital Costs Summary (USD)
Mining Capital*	$ 27,620,316
Process Capital	$ 181,571,767
Facilities Capital	$ 58,746,582
Tailings Capital	$ 21,590,251
Closure Costs	$ 1,039,987
Sub-Total	$ 290,568,903
Contingency	$ 57,360,067
Total Capital Costs	$ 347,928,970

OPEX increases from the 2022 PEA are driven by higher acid and reagent prices, diesel kiln operation and updated power tariffs.

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Table 162 Operating Costs Summary of the Lofdal PFS "Lofdal 2B-4"

Operating Costs Summary (USD)
	Life of Mine	Per tonne mined	Per tonne processed	Per kg TREO
Mining Cost	$ 652,439,644	$2.63	$37.32	$24.78
Processing	$ 996,304,067		$56.98	$37.84
G&A	$ 29,783,297		$1.70	$1.13
Total Operating Costs	$ 1,678,527,008		$96.00	$63.75

Royalties and separation costs are based on total gross revenue.

Table 163 Royalties and Separation Costs of Lofdal PFS "Lofdal 2B-4"

Other Pricing Model Costs – Life of Mine
	Base Case	Divergent Case
Royalties and Separation Costs	$ 295,383,503	$ 364,742,532

25.3. Economic Analysis

Cumulative cash flows over the 13 years mine life are: Base Case USD709,589,893 pre-tax and USD513,086,468 after-tax, and Divergent Case USD2,027,411,439 pre-tax and USD1,242,332,502 after-tax.

The Project is expected to pay back initial capital within the first 4.2 years (Base Case) and alternatively in 2.75 years (Divergent Case).

25.3.1. Sensitivity Analysis

Economic Sensitivities & Cash Flow Summary (After-Tax, 5% Discount Rate)

The Prefeasibility Study confirms that the Project delivers strong early cash flow, rapid capital recovery, and economic resilience under conservative pricing assumptions, with

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upside leverage under divergent rare earth pricing scenarios. Lofdal exhibits high sensitivity to yttrium pricing due to its HREE dominant basket.

After-Tax Cash Flow Profile

The Project achieves rapid capital payback of approximately 4.2 years after-tax in the Base Case and approximately 2.75 years in the Divergent Case, supporting strong financing attractiveness.

25.3.1.1. After-Tax Price, Cost & Exchange Rate Sensitivities

The after-tax sensitivity analysis demonstrates that:

Metal prices are the dominant value driver, with a ±20% change generating the largest impact on NPV in both, Base and Divergent cases.
Operating costs represent the second-most influential variable; however, the Project retains a positive after-tax NPV across all tested cost ranges.
Exchange rate movements provide additional economic leverage, with a weaker local currency significantly enhancing project value.
Capital costs show moderate sensitivity, confirming that the Project's value is not disproportionately dependent on Capex precision.

Process Recovery Sensitivity

Metallurgical recoveries represent a high impact, controllable value lever for the Project:

Under Base Case pricing, after-tax NPV increases from approximately USD100 million at 85% of expected recoveries to approximately USD440 million at 115% of expected recoveries.
Under Divergent Case pricing, after-tax NPV increases from approximately USD500 million at 85% of expected recoveries to approximately USD1.0 billion at 115% of expected recoveries.
The linear and consistent response of NPV to recovery improvements demonstrates that ongoing metallurgical optimisation provides meaningful value upside, while the Project remains economically viable even at materially lower-than-design recoveries.

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25.3.2. Overall Economic Interpretation

The combined cash flow and sensitivity analyses confirm that the Project:

Is financially robust under conservative assumptions;
Exhibits exceptional leverage to critical heavy rare earth pricing, particularly dysprosium, terbium and yttrium;
Benefits from strong operating margin resilience to cost pressures;
Generates early and sustained after-tax cash flow, supporting attractive project financeability; and
Provides significant embedded strategic optionality in a tightening global heavy rare earth supply environment.

25.4. Opportunities

This PFS demonstrates that the Lofdal Heavy Rare Earths 2B-4 Project has the potential to be technically and economically viable. The Project is technically uncomplicated because of the near surface nature of the deposit and relatively simple access.

Several opportunities for the project are available to further enhance the project:

Underground resource below currently planned A4;
Expansion at Area 2B pit in northeasterly direction with additional resources;
Additional potential resources in Area 5 prospect with historical HREE mineralized intercepts over 4 km strike length
Additional potential mineral resource along the regional scale mineralization trends
Destruction of lixiviant and subsequent neutralisation with Magnesium Carbonate, is costly in hydrometallurgical flowsheets. Opportunities for acid optimisation and magnesium carbonate reduction should be further investigated.

25.5. Risks

A summary of the fifteen highest initial risks with proposed mitigations are listed in Table 164.

Table 164 Project Risks

Risk Category	Risk Description	Mitigation Plan
Infrastructure	Power supply agreement and reticulation to the mine site needs cannot be achieved. Alternative power supply arrangements are required.	Combination of Own generation and Grid Power
Site Wide	Ground Conditions Impact on earthworks, foundations, and road construction quantities	Geotechnical studies required for FS study and execution.

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Risk Category	Risk Description	Mitigation Plan
Process	Ore variability affecting flotation performance	Ore Sorting included in Process Design
Tailings	Tailing characterization testwork (Geotech and Geochem) does not match design assumptions. This can impact the rate of rise, final height or liner requirement.	Schedule additional geotechnical tests inclusive of critical state analysis and shear tests during the FS. Also complete additional column kinetic tests and radionuclide tests to update the liner system requirement.
Environment	Ionizing radiation of infrastructure due to contamination during LOM.	Waste management and closure plan
Process	Sulphuric Acid quantity required neutralization and cost	Process design updated to include an acid recovery system to reduce the import of Sulphuric acid
Environment	Radon build up in confirmed spaces	Ventilation of Confined spaces. SOP for confirmed spaces and monitoring. Radiation Management plan draft in place
Process	Reagent pricing - Pricing included in the project model affects project value	Reagent pricing updated during value engineering
Process	Process reliance on expensive specialist reagents - Florea	Investigate and Test alternative suppliers/reagents
Site Wide	Availability of construction power and necessary site support services - water	Consider on site boreholes for construction. Contractors to provide their own power generators
Process	Buildup of impurities in the circuits affecting product purity and creating possible processing problems	Bleed and water treatment to manage contaminants.
Process	Process circuits adversely affected by buildup of contaminants in return water	Bleed and water treatment to manage contaminants.
Environment	Radiation contamination of process equipment cannot be adequately cleaned prior to removal from site. Interim storage is required and disposal of waste on closure is still required.	Cat 3 Storage and washing bay, Waste Management and closure plan
Environment	Safety during transportation of reagent to site and product from site	SLA with transporters should include that the supplier is responsible for safety during transportation – confirm point of transfer of liability.
Mining	Animal / human interference with trucks on the mine haul roads causes Accidents - especially at night	Consider infrared cameras on mobile equipment

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26. Recommendations

The QPs considers that additional technical work to support a Definitive Economic Assessment (DFS) is warranted for the Lofdal Project.

26.1. Metallurgical Testwork

Recommendations for further testwork:

Flotation testing on combined higher grade feed samples and upgraded low grade ore post ore sorting.
Impact of site water on flotation.
Investigate alternative fine flotation technologies.
Re-evaluate magnetic separation as a pre-concentration step considering high and low grade feed streams.
Acid recycling testing including analysis of kiln off-gas for sulphur balance.
Optimisation testing for acid bake and hydrometallurgical circuit, including testing on flotation concentrate from upgraded ore samples.

26.2. Mining

The PFS established a two-stream model as the ideal flow sheet for Lofdal, whereby higher-grade material enters flotation directly and lower-grade material will be upgraded by XRT sorting prior to flotation. Both material streams need constant supply which is challenging by operating one or two open pits only. To increase the efficiency of the upfront processing plant, the development of additional satellite pits as swing producers is advisable. These pits can supply additional supplementary high-grade or low-grade material as required during steady operation. Therefore, resource drilling of Area 5 is recommended.

Current mine life is estimated at 13 years. The PFS made provisions for an extension of mine life by designing a significantly larger than required footprint of the tailings storage facility. Inferred Resources in the northern part of Area 2B are recommended for infill drilling to potentially add additional resources at higher resource categories. Further, reconnaissance drilling at Area 5 showed significant intercepts of HREE mineralization. With its location just to the north of the planned processing plant, it is ideally situated as potential swing producer.

Mine life can likely be significantly extended by underground mining of the extension of the ore zone at Area 4 at depth. Preliminary studies showed viable options of underground mining at Area 4. These studies should be elevated to DFS level.

26.3. Mineral Processing

The following chief risks and opportunities can be addressed in future phases of the project:

Evaluate fine milling technologies as alternative to ball milling.

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Investigate opportunities to reduce the reliance on expensive specialist reagents and evaluate impact of site water on flotation chemistry.
Optimise process conditions for the hydrometallurgical circuit based on ongoing testwork using flotation concentrates produced from the upgraded sorted low grade ore samples, including:
Optimise acid bake conditions with respect to REE recovery and impurity control
Refine the acid recovery circuit including bleed requirements

The flowsheet should be updated in the next phase to reflect the optimised conditions when these are available, incorporate the improved REE precipitation circuit and optimise impurity rejection.

26.4. Infrastructure

26.4.1. Water Supply

The bulk raw water supply will in all probability be sourced from boreholes west of Fransfontein as indicated by the geohydrology study (SLR).

26.4.2. Electrical Supply

Initiate a competitive tender process for the design, supply, and construction of the 132 kV transmission line and 132/11 kV substation.

Investigate alternative electricity supply by photovoltaic.

26.4.3. Mine Access Road

For the access road construction, topographic survey of the road alignment is essential to accurately quantify the cut and fill volumes including a geotechnical assessment. Testing of existing road material or borrow pit material for suitable engineered layers recommended.

26.4.4. Tailings Storage Facility

The geochemistry analysis completed during PFS testing program indicates that a liner isn't required from an acid generation or heavy metal leaching but that the tailings indicate naturally occurring radioactive metals which are not yet fully understood in terms of risks impact to the TSF workers, groundwater and air quality. The foundation characterization would also likely induce high seepage rates to the underground and high losses in the water balance. As such, liner was adopted for the PFS as a precautionary measure, with an opportunity to reduce the lined area to the basal and embankment face areas in the next design stage.

26.5. REE Pricing

The marketing study and basket price should be reviewed and updated to reflect changing market conditions.

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26.6. Economic Analysis

Project economics may be improved by developing a full geo-metallurgy model to relate reagent consumption costs against grades, potentially Improving project economics

This PFS provides suitable economics to progress to the next stage of project development via a Definitive Feasibility study, with updated costs.

26.7. Overall

Carry out a detailed definitive feasibility study (DFS) to further develop the engineering design of the plant and recognise value engineering where possible.
Revisit the capital cost estimates in general for possible savings due to optimising the cost estimates to DFS Level.

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27. References

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Witley, J.C., Swinden, S., and Aghamirian M., 2021: NI 43-101 Technical Report – 20 May 2021 Mineral Resource Estimate, Lofdal Rare Earths Project, Namibia, prepared by MSA Group (Pty) Ltd for Namibia Critical Metals Inc., 202 pages
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28. Date and Signature Page

This report titled:

"NI 43-101 Technical Report on the Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study (PFS), Namibia" (the "Technical Report") for Namibia Critical Metals Inc.

The effective date of the report is December 3, 2025

The date of the report is January 12, 2026.

was prepared and signed by the following authors:

Signed by:

"Original Signed and Sealed"

Qualified Person

Company

Jeremy Charles Witley, Pr. Sci. Nat

The MSA Group (Pty) Ltd.

January 12, 2026

Signed by:

"Original Signed and Sealed"

Qualified Person

Company

Joseph M Keane.

SGS North America Inc.

January 12, 2026

Signed by:

"Original Signed and Sealed"

Qualified Person

Company

Peter Christians

Qubeka Mining Consultants

January 12, 2026

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Signed by:

"Original Signed and Sealed"

Qualified Person
Company
Etienne Roux
CREO Engineering Solutions (Pty) Ltd
January 12, 2026

Signed by:

"Original Signed and Sealed"

Qualified Person
Company
William van Breugel, P.Eng.
SGS Canada Inc.
January 12, 2026

Signed by:

"Original Signed and Sealed"

Qualified Person
Company
Veronique Daigle, Pr Eng.
Knight Piésold Consulting (Pty) Ltd.
January 12, 2026

SGS Bateman (Pty) Ltd

NI 43-101 Technical Report – Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study – Namibia
Page 420

29. Certificates of Qualified Persons

SGS Bateman (Pty) Ltd

NI 43-101 Technical Report – Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study – Namibia
Page 421

QP CERTIFICATE – JEREMY CHARLES WITLEY, Pr. Sci. Nat

To accompany the report entitled: NI 43-101 Technical Report on the Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study (PFS), Namibia dated January 12, 2026, and with an effective date of December 3, 2025.

I, Jeremy Charles Witley do hereby certify that:

a) I am Head of Mineral Resources at The MSA Group (Pty) Ltd, Henley House, Greenacres Office Park, Victory Park, Randburg, 2195, South Africa.
b) I graduated with a BSc (Hons) degree in Mining Geology from the University of Leicester in 1988. In addition, I obtained a Master of Science degree in Engineering from the University of Witwatersrand in 2015. I am a registered Professional Natural Scientist (Geological Science) with the South African Council for Natural Scientific Professions (SACNASP) and a Fellow of the Geological Society of South Africa. I am a "Qualified Person" for purposes of National Instrument 43-101 (the "Instrument").
c) I visited the Lofdal Property for three days from 28 to 30 October 2020, for one day on 10 November 2022 and for two days from 21 to 22 November 2023.
d) I am responsible for the preparation of Items 4, 5, 6, 7, 8, 9, 10, 11, 12, 14 and 23 of the Technical Report and co-responsible for Items 1, 2, 3, 24, 25, 26 and 27.
e) I am independent of Namibia Critical Metals as defined in Section 1.5 of National Instrument 43-101.
f) I have had no prior involvement with the subject property.
g) I have read the definition of qualified person set out in National Instrument 43-101 and certify that by virtue of my education, affiliation to a professional association, and past relevant work experience, I fulfill the requirements to be a qualified person for the purposes of National Instrument 43-101.
h) As at the effective date of the technical report, to the best of my knowledge, information and belief, this technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.
i) I have read National Instrument 43-101, Form 43-101F1 and confirm that this technical report has been prepared in accordance therewith.

Signed and dated this 12th day of January 2026.

"Original Signed and Sealed"

Jeremy Charles Witley, Pr. Sci. Nat

SGS Bateman (Pty) Ltd

NI 43-101 Technical Report – Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study – Namibia
Page 422

QP CERTIFICATE – JOSEPH M. KEANE, (B.S, M.S., PE Metallurgy)

I, Joseph M. Keane, P.E., of Tucson, Arizona, hereby certify that:

a) I am a Principal Metallurgical Engineer with SGS Minerals Services (SGS North America Inc.), with offices located at 3845 N. Business Center Drive, Suite 115, Tucson, AZ 85705.
b) I hold a B.S. in Metallurgical Engineering (Montana School of Mines, 1962) and an M.S. in Mineral Processing Engineering (Montana Tech, 1965). I am a Registered Professional Metallurgical Engineer in Arizona (12979), Colorado (17177), and Nevada (5462), and I am a Canadian National Instrument 43-101 Qualified Metallurgical Engineer (qualified to act as a "Qualified Person," as applicable, under NI 43-101). I have extensive experience providing process engineering design advice and metallurgical consulting to the mineral processing industry across a wide range of commodity projects.
c) I have not conducted a site visit of the property.
d) I am responsible for the preparation of Items 13 and 17 of the Technical Report and co-responsible for Items 1, 2, 3, 24, 25, 26 and 27.
e) I am independent of Namibia Critical Metals as defined in Section 1.5 of National Instrument 43-101.
f) I have had no prior involvement with the subject property.
g) I have read the definition of qualified person set out in National Instrument 43-101 and certify that by virtue of my education, affiliation to a professional association, and past relevant work experience, I fulfill the requirements to be a qualified person for the purposes of National Instrument 43-101.
h) As at the effective date of the technical report, to the best of my knowledge, information and belief, this technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.
i) I have read National Instrument 43-101, Form 43-101F1 and confirm that this technical report has been prepared in accordance therewith.

Signed and dated this 12th day of January 2026.

"Original Signed and Sealed"

Joseph M Keane.

SGS Bateman (Pty) Ltd

NI 43-101 Technical Report – Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study – Namibia
Page 423

QP CERTIFICATE – PETER CHRISTIANS

I, Peter Christians of Windhoek, hereby certify that:

a) I am employed as Principal Mining Engineer at Qubeka Mining Consultants Unit 2. Bougain Villas Mall, 78 Sam Nujoma Drive, Windhoek, Namibia.
b) I graduated with a Bachelor of Science Degree, Mining Engineering in 1985 from Queen's University at Kingston, Ontario Canada.
c) I am a fellow member in good standing with the Australian Institute of Mining and Metallurgy (member number 221754) since 2006. I am a "Qualified Person" for purposes of National Instrument 43-101 (the "Instrument").
d) I visited the Lofdal property on 11 December 2025.
e) I am responsible for the preparation of Items 15 and 16 of the Technical Report and co-responsible for Items 1, 2, 3, 24, 25, 26 and 27.
f) I am independent of Namibia Critical Metals as defined in Section 1.5 of National Instrument 43-101.
g) I have had no prior involvement with the subject property.
h) I have read the definition of qualified person set out in National Instrument 43-101 and certify that by virtue of my education, affiliation to a professional association, and past relevant work experience, I fulfill the requirements to be a qualified person for the purposes of National Instrument 43-101.
i) As at the effective date of the technical report, to the best of my knowledge, information and belief, this technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.
j) I have read National Instrument 43-101, Form 43-101F1 and confirm that this technical report has been prepared in accordance therewith.

Signed and dated this 12th day of January 2026.

"Original Signed and Sealed"

Peter Christians

SGS Bateman (Pty) Ltd

NI 43-101 Technical Report – Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study – Namibia
Page 424

QP CERTIFICATE – ETIENNE ALAIN ROUX, B.ENG (CHEM), SME-RM

I, Etienne Alain Roux of Windhoek, hereby certify that:

a) I am currently employed as a Director at Creo Engineering Solutions Pty Ltd, 1551 Sam Nujoma Drive, Tsumeb, Namibia.
b) I graduated with a Bachelor of Science degree in Chemical Engineering in 1999 from the University of Stellenbosch, South Africa. I am a Registered Member with the Society for Mining, Metallurgy and Exploration. I am a "Qualified Person" for purposes of National Instrument 43-101 (the "Instrument").
c) I did not visit the Lofdal Property.
d) I am responsible for the preparation of items 18 (Except Section 18.10) of the Technical Report and co-responsible for Items 1, 2, 3, 24, 25, 26 and 27.
e) I am independent of Namibia Critical Metals as defined in Section 1.5 of National Instrument 43-101.
f) I have had no prior involvement with the subject property.
g) I have read the definition of qualified person set out in National Instrument 43-101 and certify that by virtue of my education, affiliation to a professional association, and past relevant work experience, I fulfill the requirements to be a qualified person for the purposes of National Instrument 43-101.
h) As at the effective date of the technical report, to the best of my knowledge, information and belief, this technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.
i) I have read National Instrument 43-101, Form 43-101F1 and confirm that this technical report has been prepared in accordance therewith.

Signed and dated this 12th day of January 2026.

"Original Signed and Sealed"

Etienne Alain Roux, B.Eng (Chem), SME-RM

SGS Bateman (Pty) Ltd

NI 43-101 Technical Report – Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study – Namibia
Page 425

QP CERTIFICATE – WILLIAM VAN BREUGEL, P.Eng.

I, William van Breugel, P. Eng. of Saskatoon, hereby certify that:

a) I am an Associate Mining Engineer for SGS Canada Inc, with an office located at 235 Ajawan Street, Christopher Lake, Saskatchewan, Canada.

b) I graduated from the University of Waterloo in 1990 (BaSc (Hons). Geological Engineering). I am a member of good standing of the Association of Professional Engineers and Geoscientists of Saskatchewan (License #22452). I have worked as a mining engineer for over 31 years since my graduation from university. I have worked on precious metals, base metals, industrial commodities and diamond projects including mine operations and property evaluations. I am a "Qualified Person" for purposes of National Instrument 43-101 (the "Instrument").

c) I have not conducted a site visit of the property.

d) I am responsible for the preparation of Items 19, 20, 21 and 22 of the Technical Report and co-responsible for Items 1, 2, 3, 24, 25, 26 and 27.

e) I am independent of Namibia Critical Metals as defined in Section 1.5 of National Instrument 43-101.

f) I have had no prior involvement with the subject property.

g) I have read the definition of qualified person set out in National Instrument 43-101 and certify that by virtue of my education, affiliation to a professional association, and past relevant work experience, I fulfill the requirements to be a qualified person for the purposes of National Instrument 43-101.

h) As at the effective date of the technical report, to the best of my knowledge, information and belief, this technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

i) I have read National Instrument 43-101, Form 43-101F1 and confirm that this technical report has been prepared in accordance therewith.

Signed and dated this 12th day of January 2026.

"Original Signed and Sealed"

William van Breugel, P.Eng.

SGS Bateman (Pty) Ltd

NI 43-101 Technical Report – Lofdal Heavy Rare Earths Project 2B-4 Preliminary Feasibility Study – Namibia
Page 426

QP CERTIFICATE – VERONIQUE DAIGLE, Pr. Eng and Eng.

I, Veronique Daigle of Windhoek, hereby certify that:

a) I am employed as Lead Engineer and a Director at Knight Piésold Consulting (Pty) Ltd (registration 2008:0657), at 11 Nelson Mandela, Klein Windhoek, Windhoek, Namibia.
b) I graduated with a Civil Engineering Degree (Cooperative Program) in 2006 from the Université de Sherbrooke, in the Province of Québec, Canada.
c) I am a member in good standing with the Engineering Council of Namibia and Registered as Professional Engineer (license number PE2017-19) since 2017. My Relevant Experience includes 18 years of continuous experience in tailings, geotechnical engineering and water management employed at Knight Piésold. I am a member in good standing with the South African Committee on Large Dams, the Canadian Dam Association, and the Ordre des Ingenieurs du Québec (License No 143 749), Canada. I am a "Qualified Person" for purposes of National Instrument 43-101 (the "Instrument").
d) I have not conducted a site visit of the property.
e) I am responsible for the preparation of Items 18.10 of the Technical Report and co-responsible for Items 1, 2, 3, 24, 25, 26 and 27.
f) I am independent of Namibia Critical Metals as defined in Section 1.5 of National Instrument 43-101.
g) I have had no prior involvement with the subject property.
h) I have read the definition of qualified person set out in National Instrument 43-101 and certify that by virtue of my education, affiliation to a professional association, and past relevant work experience, I fulfill the requirements to be a qualified person for the purposes of National Instrument 43-101.
i) As at the effective date of the technical report, to the best of my knowledge, information and belief, this technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.
j) I have read National Instrument 43-101, Form 43-101F1 and confirm that this technical report has been prepared in accordance therewith.

Signed and dated this 12th day of January 2026.

"Original Signed and Sealed"

Veronique Daigle, Pr. Eng., Director at Knight Piésold Consulting (Pty) Ltd

SGS Bateman (Pty) Ltd

	La	Ce	Pr	Nd	Sm	Eu	Gd	Tb	Dy	Ho	Er	Tm	Yb	Lu	Y
10%	76	77	78	82	77	80	80	79	82	81	80	79	82	80	83
20%	94	95	95	96	97	97	97	97	96	96	97	96	96	96	96

	La	Ce	Pr	Nd	Sm	Eu	Gd	Tb	Dy	Ho	Er	Tm	Yb	Lu	Y
10%	76	77	78	82	77	80	80	79	82	81	80	79	82	80	83
20%	94	95	95	96	97	97	97	97	96	96	97	96	96	96	96

AI assistant

Regency Silver Corp. — Audit Report / Information 2025

35757_rns_2026-01-13_ae982f37-1712-4261-b855-dbe9962e8937.pdf

LOFDAL HEAVY RARE EARTHS PROJECT 2B-4 PRELIMINARY FEASABILITY STUDY (PFS) NAMIBIA

Qualified Persons

Company

TABLE OF CONTENTS

LIST OF FIGURES

LIST OF TABLES

1. Summary

1.1. Introduction

1.2. Accessibility, Climate, Local Resources, Infrastructures, and Physiography

1.3. Geology and Mineralization

1.4. Exploration Status

Starter pit at Area 4 and bulk sample extraction

1.5. Mineral Processing and Metallurgical Testing

Ore Sorting

Flotation test program

Hydrometallurgical test work and results

1.6. Mineral Resource Estimate

1.7. Mineral Reserve Estimate

1.8. Mining Methods

1.9. Recovery Methods

1.10. Project Infrastructure

1.11. Rare Earth Pricing

Divergence price trends

1.12. Socio-Economic and Environmental Impact

1.13. Legal and Statutory

1.14. Capital and Operating Cost Estimate

1.15. Economic Analysis

Sensitivity Analysis

Economic Sensitivities & Cash Flow Summary (After-Tax, 5% Discount Rate)

After-Tax Cash Flow Profile

After-Tax Price, Cost & Exchange Rate Sensitivities

Process Recovery Sensitivity

Overall Economic Interpretation

1.16. Conclusions

1.17. Recommendations

2. Introduction

2.1. Purpose of Report

2.2. Terms of Reference

2.3. Qualifications of Consultants

2.4. Report Responsibility and Qualified Persons

2.5. Site Visit

2.6. Currency, Units, Abbreviations and Definitions

2.7. Effective Date

2.8. Previous Technical Reports

3. Reliance on other Experts

4. Property Description and Location

4.1. Property Location

4.2. Property Description

4.2.1. Mining License (ML) 200

Part 2 – Work Program and Obligations

Part 3 – Environment

Part 5 – Additional Conditions

4.2.2. General Provisions

4.2.3. Adjacent and Overlapping EPLs

5. Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1. Accessibility

5.2. Climate

5.3. Local Resources and Infrastructure

5.4. Physiography

6. History

7. Geological Setting and Mineralization

7.1. Regional Geology

7.2. Local Geology

7.2.1. The Huab Metamorphic Complex

7.2.2. The Damara Orogen

7.2.3. Early Damaran Alkaline / Carbonatitic Intrusions

7.2.3.1. The Oas Syenite

7.2.3.2. The Lofdal Carbonatite Complex

7.2.3.2.1. Nepheline Syenite

7.2.3.2.2. Phonolite

7.2.3.2.3. Lofdal Breccias

7.2.3.2.4. Carbonatites

7.3. Structural Setting

7.4. REE Mineralization

7.4.1. Regional Setting

7.4.2. Mineralization in Area 4

7.4.3. Mineralization in Area 2B

Regency Silver Corp. Audit Report / Information 2025

	La	Ce	Pr	Nd	Sm	Eu	Gd	Tb	Dy	Ho	Er	Tm	Yb	Lu	Y
10%	76	77	78	82	77	80	80	79	82	81	80	79	82	80	83
20%	94	95	95	96	97	97	97	97	96	96	97	96	96	96	96