ROKMASTER RESOURCES CORP.

NI 43-101 TECHNICAL REPORT

AN UPDATED PRELIMINARY ECONOMIC ASSESSMENT OF THE

REVEL RIDGE PROJECT, REVELSTOKE, B.C., CANADA

Effective Date of Resource Estimate: 29 January, 2020 Effective Date of Preliminary Economic Assessment: 08 December, 2020 Report Signature Date: 22 January, 2021

Report By

Eugene Puritch, P.Eng., FEC, CET Fred Brown, P.Geo. Jarita Barry, P.Geo. Richard Routledge, P.Geo. Nigel Fung, P.Eng. Richard M. Gowans, P.Eng.

1.0	SUMMARY1
2.0	INTRODUCTION AND TERMS OF REFERENCE7
2.1	SOURCES OF INFORMATION8
2.2	UNITS AND CURRENCY8
2.3	GLOSSARYAND ABBREVIATION OF TERMS8

3.0	RELIANCE ON OTHER EXPERTS12
4.0	PROPERTY DESCRIPTION AND LOCATION13
4.1	PROPERTY LOCATION13
4.2	HUAKAN –ROKMASTER AGREEMENT TERMS13
4.3	PROPERTY DESCRIPTION14
4.4	GENERAL REQUIREMENTS FOR MINERAL CLAIMS18
4.5	PERMITTING18
4.6	ARMEX STATEMENT OF CLAIM18
4.7	FIRST NATIONS WITH POTENTIAL INTERESTS IN THE REVELSTOKE
	REGION18

5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE
	AND PHYSIOGRAPHY20
5.1	ACCESSIBILITY20
5.2	CLIMATE20
5.3	LOCAL RESOURCES20
5.4	INFRASTRUCTURE21
5.5	PHYSIOGRAPHY24

6.0	HISTORY26
6.1	PREVIOUS MINERAL RESOURCE ESTIMATES30
7.0	GEOLOGICAL SETTING AND MINERALIZATION32
7.1	REGIONAL GEOLOGY32
7.2	PROPERTY GEOLOGY32
	7.2.1HamillGroup32
	7.2.2Mohican Formation32
	7.2.3Badshot Formation36
7.3	LARDEAU GROUP36
	7.3.1Index Formation36
	7.3.2Micaceous Quartzite Unit36
	7.3.3Jowett Formation36
7.4	LOCAL GEOLOGY37
7.5	STRUCTURE37
7.6	MINERALIZATION38
	7.6.1Main Zone38

	7.6.2Hanging Wall and Footwall Zones40
	7.6.3Yellowjacket Zone40
	7.6.4Other Showings41

8.0	DEPOSIT TYPES43
8.1	MAIN ZONE43
8.2	YELLOWJACKET ZONE44
9.0	EXPLORATION45
10.0	DRILLING46
10.1	HUAKAN DIAMOND DRILLING46
	10.1.1Collar Surveying51
	10.1.2Downhole Surveying51
	10.1.3Core Recovery and Storage51
	10.1.4Drill Core Size and Orientation52

	10.1.5Contractor52
	10.1.6Comments52
11.0	SAMPLE PREPARATION, ANALYSES AND SECURITY53
11.1	SAMPLING METHOD AND APPROACH BY HUAKAN53
11.2	CHAIN OF CUSTODY54
11.3	SAMPLE PREPARATION AND ANALYSES54
11.4	HUAKAN QUALITY ASSURANCE/QUALITY CONTROL55
	11.4.1Certified Reference Material55

12.0	DATA VERIFICATION58
12.1	2010/2011 SITE VISIT AND INDEPENDENT SAMPLING58
12.2	2012 SITE VISIT AND INDEPENDENT SAMPLING61
12.3	DATABASE VERIFICATION64

13.0	MINERAL PROCESSING AND METALLURGICAL TESTING	65
13.1	INTRODUCTION65
	13.1.1Historical Testwork65
	13.1.2Recent Mineral Processing Testwork (2011-2013)	66
	13.1.3Recent Flowsheet Development Testwork (2020)77
	13.1.4Recent Precious Metals Extraction Testwork (2011-2013)78
13.2	RECOMMENDATIONS80

14.0	MINERAL RESOURCE ESTIMATE82
14.1	INTRODUCTION82
14.2	DATA SUPPLIED82
14.3	DATABASE VALIDATION83
14.4	BULK DENSITY83
14.5	ECONOMIC PARAMETERS83

14.7	COMPOSITING85
14.8	EXPLORATORY DATA ANALYSIS	86
14.9	TREATMENT OF EXTREME VALUES	89
14.10	CONTINUITY ANALYSIS90
14.11	BLOCK MODEL90
14.12	ESTIMATIONAND CLASSIFICATION91
14.13	MINERAL RESOURCE ESTIMATE93
14.14	VALIDATION94

15.0	MINERAL RESERVE ESTIMATE104
16.0	MINING METHODS105
16.1	SUMMARY105
16.2	MINE PLAN110
16.3	DEVELOPMENT LAYOUT114
16.4	STOPING117
16.4.1	Main Zone119
16.4.2	Yellowjacket Zone120
16.5	EQUIPMENT REQUIREMENTS	121
16.6	UNDERGROUND OPERATIONS PERSONNEL124
16.7	VENTILATION125
16.8	SCHEDULE126

17.0	RECOVERY METHODS129
17.1	PROCESS DESIGN CRITERIA129
17.1.1	Design Basis129
17.1.2	Process Design Criteria132
17.2	PROCESS DESCRIPTION138
17.2.1	Crushing138
17.2.2	Heavy Media Separation138
17.2.3	Grinding139
17.2.4	Gravity Separation Circuit	139
17.2.5	Lead Flotation139
17.2.6	Zinc Flotation140
17.2.7	Bulk Sulphide-Gold Flotation140
17.2.8	Dewatering140
17.2.9	Pressure Oxidation141
	17.2.10Gold Leaching and Recovery141
	17.2.11Reagents141
	17.2.12Services142
17.3	PROCESSING PLANT LAYOUT142
18.0	PROJECT INFRASTRUCTURE144
18.1	ACCESS144
18.2	POWER SUPPLY144
18.3	WATER SYSTEMS144

18.3.1	Process Water144
18.3.2	Fresh Water144
18.3.3	Potable Water144
18.4	FUEL STORAGE145
18.5	BUILDINGS145
18.6	SEWAGE TREATMENT145
18.7	FIRE PROTECTION145
18.8	TAILINGS DISPOSAL145
18.9	VENTILATION145
18.10	WASTE DUMP146
18.11	CEMENTED BACKFILL PLANT146
18.12	EXPLOSIVES STORAGE146
18.13	MINE DEWATERING AND SEDIMENTATION PONDS147
18.14	MAINTENANCE SHOPS, LAYDOWN AND STAGING AREA147
18.15	WASTE MANAGEMENT FACILITY147
18.15.1	Basis of Design147
18.15.2	Waste Management Facility Layout148
18.15.3	Waste Management Facility Construction150
18.15.4	Water Management151
18.15.5	Reclamation and Closure152
18.15.6	Material, Quantities and Cost Estimates152
18.15.7	Recommendations/Potential Opportunities153


19.0	MARKET STUDIES AND CONTRACTS154

20.0	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR
	COMMUNITY IMPACT155
20.1	ENVIRONMENTAL STUDIES AND POTENTIAL ISSUES155
20.1.1	Overview155
20.1.2	Climate and Meteorology155
20.1.3	Surface Water Quantity and Quality156
20.1.4	Groundwater Quality and Quantity156
20.1.5	Fisheries and Fish Habitat156
20.1.6	Vegetation, Forestry and Wetlands156
20.1.7	Wildlife and Wildlife Habitat157
20.2	WASTE MANAGEMENT157
20.3	WATER MANAGEMENT158
20.3.1	Water Balance158
20.3.2	Water Treatment Conceptual Basis161
20.3.3	Water Treatment Conceptual Design161
20.4	PERMITTING REQUIREMENTS164
20.5	ENVIRONMENTAL, SOCIAL AND HEALTH MANAGEMENT SYSTEM166
20.620.7	SOCIAL AND COMMUNITY ISSUES167FIRST NATIONS168

21.0		CAPITAL AND OPERATING COSTS170
21.1		CAPITAL COSTS170
	21.1.1	Mining Capital171
	21.1.2	Processing Capital172
	21.1.3	Infrastructural Capital173
	21.1.4	Owner's Costs174
	21.1.5	Mine Closure Costs174
	21.1.6	Contingency174
21.2		OPERATINGCOSTS174
	21.2.2	Mine Operating Costs175
	21.2.3	Processing Operating Costs176
	21.2.4	General and Administrative Costs177
	21.2.5	Selling Costs177
22.0		ECONOMIC ANALYSIS178
22.1		CAUTIONARY STATEMENT178
22.2		BASIS OF EVALUATION179
22.3		MACRO-ECONOMIC ASSUMPTIONS179
	22.3.1	Exchange Rate and Inflation179
	22.3.2	Weighted Average Cost of Capital179
	22.3.3	Expected Metal Prices180
	22.3.4	Taxation Regime180
	22.3.5	Royalty180
22.4		TECHNICAL ASSUMPTIONS181
	22.4.1	Product Offtake181
	22.4.2	Production Schedule181
22.5		OPERATING MARGIN (BASE CASE)182
22.6		PROJECT CASH FLOW (BASE CASE)182
22.7		SENSITIVITY STUDY AND RISK ANALYSIS184
	22.7.1	Metal Price and Exchange Rate Assumptions184
	22.7.2	Alternative Development Case (Toll Treatment)	185
22.8		CONCLUSION188

23.0		ADJACENT PROPERTIES189
24.0		OTHER RELEVANT DATA AND INFORMATION190
25.0		INTERPRETATION AND CONCLUSIONS191

26.0		RECOMMENDATIONS193
27.0		DATE AND SIGNATURE PAGE194
28.0		REFERENCES195
29.0		CERTIFICATES202

List of Tables

Table 1.1	(1-7)Revel Ridge 2020 Mineral Resource Estimate3
Table 1.2	Revel Ridge 2020 PEA Detailed Parameters and Outputs4
Table 1.3	After-Tax NPV and IRR Sensitivities to Commodity Prices6
Table 1.4	Budget for Proposed PFS Program6
Table 2.1	List of Abbreviations8
Table 4.1	Revel Ridge Mineral Claims15
Table 4.2	Revel Ridge Crown Grant Lots15
Table 6.1	1996 Yellowjacket Zone Drill Highlights27
Table 6.2	2007 Yellowjacket Zone Drill Highlights28
Table 6.3	Previous 2011 Mineral Resource Estimate30
Table 6.4	Previous 2012 Mineral Resource Estimate31
Table 6.5	Previous 2018 Mineral Resource Estimate31
Table 10.1	Drill Programs Summary46
Table 10.2	Drill Programs Summary Main Zone Drill Highlights in Area A	47
Table 10.3	Main Zone Drill Highlights in Area B48
Table 10.4	Drill Programs Summary Main Zone Drill Highlights in Area C49
Table 10.5	Drill Programs Summary Main Zone Drill Highlights in Area D	49
Table 10.6	Drill Programs Summary Main Zone Drill Highlights in Area E50
Table 10.7	Drill Programs Summary Main Zone Drill Highlights in Area F	50
Table 10.8	Drill Programs Summary Yellowjacket Zone Drill Highlights in AreaF50
Table 13.1	Main Zone Metallurgical Composite Sample Chemical Analyses67
Table 13.2	Yellow Jacket Metallurgical Sample Chemical Analyses67
Table 13.3	Bond Ball Mill Index Test Results69
Table 13.4	Heavy Liquid Test Results –Main Zone69
Table 13.5	Summary of Flotation Test Results –Main Zone Composite JL1LCT172
Table 13.6	Summary of Flotation Test Results –Main Zone Composite JL1LCT272
Table 13.7	Summary of the Main Zone Variability LCT Results73
Table 13.8	Locked Cycle Float Tests –Main Zone Calculated Heads73

Table 13.9	Summary of Yellowjacket Open Circuit Flotation Cleaner TestResults74
Table 13.10	Summary of Flotation Test Results –Yellowjacket Composite 5LCT175
Table 13.11	Final Flotation Concentrate Multi-Analyses for Main Zone Test FLC1and Yellowjacket FLC176
Table 13.12	Summary of POX Testwork Results Including Precious MetalsCyanide Leach Extractions80
Table 14.1	Database Summary82
Table 14.2	Bulk Density Values83
Table 14.3	Revel Ridge Economic Parameters –Main Zone84
Table 14.4	Revel Ridge Economic Parameters –Yellowjacket Zone84
Table 14.5	Composite Summary Statistics87
Table 14.6	Main Zone Composite Correlation Matrix87
Table 14.7	Capping Thresholds89
Table 14.8	Block Model Setup90
Table 14.9	Search Parameters91
Table 14.10	(1-7)Revel Ridge 2020 Mineral Resource Estimate93
Table 14.11	Main Zone Validation Statistics94
Table 16.1	Stope Shape Tonnes and Grades: Main Zone109
Table 16.2	Stope Shape Tonnes and Grades: Yellowjacket Zone110
Table 16.3	Development Plan for Stope Levels113
Table 16.4	Development Metres by Level and Type115
Table 16.5	Undiluted Main Zone Mineral Resources included in the Mine Plan119
Table 16.6	Diluted Mill Feed from the Main Zone Mineral included in the MinePlan120
Table 16.7	Undiluted Yellowjacket Zone Mineral Resources included in the MinePlan120
Table 16.8	Diluted Mill Feed from the Yellowjacket Zoneincluded in the MinePlan120
Table 16.9	Summary of Underground Operations Personnel125
Table 16.10	Mine Ventilation126
Table 16.11	Life-of-Mine Production Schedule128
Table 17.1	Process Design Basis131

Page

Table 17.2	Process Design Criteria –Mineral Processing133
Table 17.3	Process Design Criteria –Gold Recovery Plant135
Table 18.1	Waste Management Facility Key Design Criteria148
Table 20.1	Estimated Major Sources and Sinks of Water160
Table 20.2	First Nations with Potential Interest in the Project Area168
Table 21.1	Capital Cost Summary –Base Case170
Table 21.2	Capital Cost Summary –Alternative Case171
Table 21.3	Capital Cost Summary –Mining171
Table 21.4	Initial Capital Cost Summary –Processing172
Table 21.5	Initial Capital Cost Summary –Indirect173
Table 21.6	Initial Capital Cost Summary –Infrastructure173
Table 21.7	Initial Capital Cost Summary –Owner's Costs174
Table 21.8	LOM Total Cash Costs –Base Case175
Table 21.9	LOM Total Cash Costs –Alternative Case175
Table 21.10	Mine Operating Costs (Base Case)175
Table 21.11	Process Operating Costs (Base Case)176
Table 21.12	Process Operating Costs (Alternative Case)177
Table 22.1	Product Offtake Terms181
Table 22.2	Life-of-Mine Cash Flow Summary –Base Case182
Table 22.3	Base Case Life of Mine Annual Cash Flow183
Table 22.4	Base Case: Sensitivity of NPV, IRR and Payback to Metal Price185
Table 22.5	Toll Treatment: Life-of-Mine Cash Flow Summary186
Table 26.1	Budget for Proposed PFS Program193

List of Figures

Figure 4.1	Property Location Map13
Figure 4.2	Regional Location Map16
Figure 4.3	Property Boundary Map17
Figure 5.1	3-D View of Underground Workings22
Figure 5.2	Plan View of the Underground Workings of 832 m and 830 m LevelDrifts22
Figure 5.3	Plan View of Roads, Camp, Shop, PAG Pile and Projection of 832 mand 830 m Level Underground Workings24
Figure 7.1	Stratigraphic Column33
Figure 7.2	Revel Ridge Regional Geology Map34
Figure 7.3	Revel Ridge Site Geology Map35
Figure 7.4	Plan showing Surface Traces of Mineralized Zones42
Figure 10.1	Longitudinal Section Showing All Drill Hole Pierce Points and 2012Drilling Areas48
Figure 12.1	2010/2011 P&E Verification Samples for Gold59
Figure 12.2	2010/2011 P&E Verification Samples for Silver59
Figure 12.3	2010/2011 P&E Verification Samples for Lead60
Figure 12.4	2010/2011 P&E Verification Samples for Zinc60
Figure 12.5	2012 P&E Verification Samples for Gold62
Figure 12.6	2012 P&E Verification Samples for Silver62
Figure 12.7	2012 P&E Verification Samples for Lead63
Figure 12.8	2012 P&E Verification Samples for Zinc63
Figure 13.1	Main Zone Heavy Liquid Test Results for Different SampleComposites70
Figure 13.2	Main Zone Locked Cycle Flotation Test FLC1 Flowsheet71
Figure 13.3	Yellowjacket Zone Locked Cycle Flotation Test FLC1 Flowsheet75
Figure 13.4	Gold Gravity Concentrate Grade vs Recovery Curve78
Figure 14.1	Isometric Projection of Mineral Resource Domains85
Figure 14.2	Dotplot of Constrained Assay Lengths86
Figure 14.3	Main Zone QQ Plots for Drill Hole vs. Chip Composites88
Figure 14.4	Capping Analysis Plots89
Figure 14.5	Isometric Projection of Block Classification (Looking West)93

Page

Figure 14.6	Main Zone Swath Plots95
Figure 14.7	Drill Hole Plan and Vertical Cross-Sections98
Figure 16.1	2020 Block Model and Stope Shapes: Plan View105
Figure 16.2	2020 Block Model and Stope Shapes: Section View106
Figure 16.3	Main Zone Stopes Included in the Mine Plan Coded for NSR Values106
Figure 16.4	Yellowjacket Zone Stopes Included in the Mine Plan Coded for NSRValues107
Figure 16.5	Photo of 832 m Elevation Portal in September, 2020108
Figure 16.6	Schematic Cross-and Longitudinal Section of Mining Method109
Figure 16.7	Schematic Showing Conceptual Plan for Waste Development in theLower Mine (Not to Scale)117
Figure 16.8	Insulated Shop123
Figure 17.1	Main Zone Mineral Processing Block Flow Diagram130
Figure 17.2	Main Zone Gold Recovery Block Flow Diagram131
Figure 17.3	Proposed Process Flow Diagram137
Figure 17.4	Proposed Process Plant Layout143
Figure 18.1	A Plan View and Cross-Section of the Conceptual WMF149
Figure 20.1	Conceptual Project Water Balance160
Figure 22.1	Ten Year Metal Price History180
Figure 22.2	LOM Production Schedule181
Figure 22.3	LOM Net Revenue and Operating Costs (Base Case)182
Figure 22.4	Life-of-Mine Base Case Cash Flows184
Figure 22.5	Sensitivity of Base Case NPV7.5to Capital, Operating Costs and MetalPrices185
Figure 22.6	Life-of-Mine Cash Flows (Alternative Case)187
Figure 22.7	Sensitivity of NPV7.5in Alternative Case187

1.0 SUMMARY

The following report was prepared to provide a National Instrument (NI) 43-101 Updated Technical Report on the Revel Ridge Property (the Property), formerly named the J&L property, for Rokmaster Resources Corp. ("Rokmaster" or the "Company"). This Technical Report discloses the results of a Preliminary Economic Assessment (PEA) of the Property prepared by Micon International Limited (Micon) based on a Mineral Resource Estimate prepared by P&E Mining Consultants Inc. (P&E) that has an effective date of January 29, 2020. Rokmaster is a British Columbia corporation trading on the TSX Venture Exchange with the symbol RKR and on the OTCQB with the symbol RKMSF.

The Property hosts two known and significant polymetallic precious and base metal deposits, the Main Zone and the Yellowjacket Zone, which are located 35 kilometres north of Revelstoke, British Columbia, Canada. The Property consists of 18 mineral tenure claims and 10 Crown Grant Lots for a total of 3,150.74 hectares.

Rokmaster has an option agreement dated December 23, 2019 to earn a 100% interest in the Property from Huakan International Mining Inc. (Huakan), formerly Merit Mining Corp. (Merit). The agreement provides for Rokmaster to earn a 100% interest in the Property and associated assets without any underlying royalties. Rokmaster has been advised that a legal action has arisen between Armex Mining Corp. (Armex) and Huakan whereby Armex claims that it has a valid letter of intent with Huakan covering the Property. Huakan has notified Armex that it intends to defend the Armex action and has filed a counter claim against Armex. The legal action has not been resolved at the time of this Technical Report, but Huakan has fully indemnified Rokmaster from any potential losses.

The Property lies within the Selkirk Mountains near the north end of the Kootenay Arc, a complex sequence of east dipping Neoproterozoic to Lower Paleozoic metasedimentary and metavolcanic miogeosynclinal rocks. The belt is characterized by tight to isoclinal folds and generally west verging thrust faults with greenschist grade regional metamorphism. The Revel Ridge Property is underlain by north to northwest striking, moderate to steeply east dipping metasediments and metavolcanic rocks of the Hamill and Lardeau Group and Badshot and Mohican Formation rocks.

The Main Zone is a structurally controlled orogenic gold deposit comprising polymetallic massive sulphides containing gold, silver, lead and zinc. The Main Zone is a sheet-like tabular sulphide vein system hosted in a large planer deformation zone composed of banded massive and stringer arsenopyrite-pyrite-sphalerite-galena mineralization with appreciable content of gold and silver. The Main Zone has been traced on surface by prospecting, trenching and soil sampling for a strike length of over 3 kilometres. Drilling has intersected the zone over a 1,500 metre strike length and 800 metres down dip. The Main Zone generally dips at approximately 56 degrees to the northeast with an average true thickness of 2.5 metres, however, it can reach 15 metres in true thickness and has the potential to be expanded beyond the current drilled limits.

The silver-zinc-lead rich Yellowjacket Zone is considered to be a structurally-controlled carbonate-hosted replacement deposit composed of multiple parallel siliceous sphaleritegalena-bearing zones. The individual zones making up the Yellowjacket Zone occur as lenticular bodies each up to 8 metres thick at the contact between alternating units of quartzites, phyllitic sediments and limestone. Currently, the Yellowjacket Zone has not been shown to be as laterally extensive as the Main Zone but, pending further exploration, it remains open to the northwest and down plunge. The Yellowjacket Zone sub parallels and is in the immediate hanging wall of the Main Zone. The Yellowjacket Zone has little notable gold, however, it has higher silver, lead and zinc values than the Main Zone from which it is metallurgically distinct. In contrast to better known Kootenay Arc silver-lead-zinc occurrences having grades of 2–4 g/t silver, grades at the Yellowjacket zone are much higher, 62.6 g/t Ag (Puritch et al., 2020).

Numerous exploration companies including several major mining companies have explored and advanced the Property since the Main Zone's discovery in 1912. At least 315 diamond drill holes have been completed on the Property from 1983 to present, totalling 41,075.9 metres of drilling. A total of 3.1 kilometres of underground workings are present on the Property. A 1.4-kilometre-long track drift (2.4 m x 2.4 m profile) at the 830 m level has exposed the Main Zone for approximately 800 metres in length. The 550-metre-long (5 m x 5 m profile) 832 m level trackless drift installed by Merit in 2008, connects to the 830 m track drift, providing underground access to the 830 m drift. Five crosscuts totalling 1,150 metres provided access to drill stations that were utilized to drill-define the deposits. Several raises have aided in the extraction of several bulk samples. There is an adit and accompanying drift extending 152 metres along the Main Zone called the "986 m level" that is now inaccessible.

In late 2010, Merit/Huakan completed a 60-hole, 7,897 metre underground drill program focused on the Main Zone. This program had the objective of verifying historic drilling and sampling and infilling an 800-metre strike by 200 metre dip of the Main Zone with 30 metre drill centres. This program led to P&E completing the first NI 43-101 Mineral Resource Estimate on the Property in September, 2011 and a subsequent PEA by Micon in May, 2012 based on the 2011 Mineral Resource Estimate.

The 2010 exploration program was followed in 2012 by a 450-metre drifting and a 45-hole, 9,725 metre underground drill program to expand the Mineral Resource Estimate of the Main Zone. The 2012 program was successful in increasing the Mineral Resources. Results of an Updated Mineral Resource Estimate by P&E were reported in a news release by Huakan dated September 18, 2012. This estimate significantly increased Indicated Mineral Resources on the Main Zone and for the first time included a Mineral Resource Estimate on the Yellowjacket Zone. No subsequent material physical work has been done on the Property since the 2012 Updated Resource Estimate. In January, 2013, Huakan reported updated metallurgical test work results from a bulk sample collected in the 2012 program.

In its January, 2020 Technical Report, P&E updated the Revel Ridge Mineral Resource Estimate to include current trailing metal prices, mining costs, and exchange rates as well as updated metallurgical test results (see Table 1.1).

MineralizedZone	Classification	Tonnes(k)	Au(g/t)	Au(koz)	Ag(g/t)	Ag(koz)	Pb(%)	Zn(%)	Au Eq(g/t)	Au Eq(koz)
Main Zone	Measured	1,352	6.13	266	62.8	2,730	2.19	4.09	9.14	397
	Indicated	2,848	5.33	488	49	4,487	1.72	3.11	7.56	692
	Meas & Ind	4,200	5.59	755	53.4	7,216	1.87	3.43	8.07	1,089
	Inferred	4,562	4.36	639	61.8	9,064	1.88	2.59	6.55	961
	Indicated	298	0.91	9	55.3	530	2.50	5.72	4.70	45
HW Zone	Inferred	38	0.22	0	75	92	3.08	5.44	4.34	5
FW Zone	Inferred	341	3.91	43	25.3	277	0.53	0.48	4.20	46
Yellowjacket	Indicated	771	0.09	2	62.6	1,552	2.60	9.93	NA	NA
Zone	Inferred	23	0.11	0	55.4	41	2.65	7.68	NA	NA

Table 1.1 Revel Ridge 2020 Mineral Resource Estimate (1-7)

Mineral Resources which are not Mineral Reserves do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues.
The Inferred Mineral Resource in this estimate has a lower level of confidence than that applied to an Indicated Mineral Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of the Inferred Mineral Resource could be upgraded to an Indicated Mineral Resource with continued exploration.
The Mineral Resources in this estimate were calculated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.
The following parameters were used to derive the NSR block model cut-off values used to define the Mineral Resource: Dec 31, 2019 US$ two-year trailing avg. metal prices:

Pb $0.96/lb, Zn $1.24/lb, Au $1,331/oz, Ag $15.95/oz
Exchange rate of US$0.76 = CDN $1.00
Process recoveries of Pb 74%, Zn 75%, Au 91%, Ag 80%
Smelter payables of Pb 95%, Zn 85%, Au 96%, Ag 91%
Refining charges of Au US$10/oz, Ag US$0.50/oz
Concentrate freight charges of CDN$65/t and Smelter treatment charge of US$185/t
Mass pull of 5% and 8% concentrate moisture content.
1. NSR cut-off of CDN$110 per tonne was derived from $75/t mining, $25/t processing, $10/t G&A.
1. AuEq = Au g/t + (Ag g/t x 0.011) + (Pb % x 0.422) + (Zn % x 0.455)
1. Above parameters derived from 2012 PEA and other similar benchmarked projects.

Both the Main Zone and the Yellowjacket Zone have potential for further expansion. The Main Zone, in particular, remains open in a number of directions. It has a tabular, predictable geometry and grade distribution and is laterally extensive as defined by drilling to date. Its surface strike length has been established to be in excess of three kilometres, of which only a portion has been drill-tested.

The 2020 Revel Ridge PEA considers an underground mine with on-site treatment of the mined material by conventional milling, gravity and flotation to produce concentrates for sale to thirdparty smelters, in combination with on-site treatment of refractory gold concentrates to produce gold-silver doré. The mine will comprise an owner-operated, ramp developed, long hole stope underground mine.

The process plant capacity of 2,300 tonnes per day will result in a production lifespan of 12 years. An additional 18 months of mine ramp access and development, and construction of the process plant and dry-stack tailings facility is planned prior to the project becoming fully operational in Year 1. The PEA leverages Revel Ridge's extensive existing infrastructure,

including all-weather access roads, local hydroelectric facilities, 3 km of underground development, permitted waste rock storage facility, full camp facility and proximity to the City of Revelstoke with its skilled labour pool.

This PEA is derived from the Company's NI 43-101 Mineral Resource Estimate (January 29, 2020), and does not include results from the recently initiated and ongoing 2020 Phase I exploration diamond drilling program. The effective date of the PEA is December 08, 2020.

Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. The PEA is preliminary in nature and includes Inferred Mineral Resources that are too speculative to have economic considerations applied to them that would enable them to be classified as Mineral Reserves. There is no certainty that PEA results will be realized.

Assumptions
Gold Price (US$/oz)	$1,561
Silver Price (US$/oz)	$20.55
Zinc Price (US$/lb)	$1.07
Lead Price (US$/lb)	$0.91
Exchange Rate (US$/CDN$)	0.77
Royalties	0%
Contained Metals in Process PlantFeed
Contained Gold Ounces (koz)	1,280
Contained Silver Ounces (koz)	15,934
Contained AuEq Ounces (koz)	1,785
Mining
Mine Life (Years)	12
Main Zone LOM production (Mt, diluted)	9.39
Average Diluted Gold Grade (g/t)	4.24
Average Diluted Silver Grade (g/t)	49.8
Average Diluted Zinc Grade %	2.62
Average Diluted Lead Grade %	1.63
Yellow Jacket Zone LOM production (Mt, diluted)	0.65
Average Diluted Gold Grade (g/t)	0.06
Average Diluted Silver Grade (g/t)	43.00
Average Diluted Zinc Grade %	7.47
Average Diluted Lead Grade %	1.90
Processing
Processing Throughput (tpd)	2,300
Total Mill-feed Tonnage (Mt)	10,04
Average revenue per tonne treated (CDN$/t)	300.25
Average Diluted Gold Equivalent Grade (g/tAu Eq)	5.53
Gold Recovery (overall)	83.5%
Silver Recovery (overall)	52.0%

Table 1.2 Revel Ridge 2020 PEA Detailed Parameters and Outputs

Production
LOM Gold Production (koz)	1,068
LOM Silver Production (koz)	8,282
LOM Zinc Production (Mlbs)	450
LOM Lead Production (Mlbs)	255
LOM Gold Equivalent Production (koz Au Eq)	1,490
LOM Average Annual Gold Production (koz)	89
LOM Average Annual Silver Production (koz)	690
LOM Average Annual Gold Equivalent Production (koz)	124
Operating Costs
Mining Cost (CDN$/tMilled)	$62.42
Processing Cost (CDN$/tMilled)	$65.07
G&A Cost (CDN$/tMilled)	$7.59
Total Operating Cost (CDN$/tMilled)	$135.08
Cash Costs and AISC
LOM Cash Cost (US$/oz Au) Net of Silver-Zinc-Lead By-Products	$362
LOM Cash Cost (US$/oz AuEq) Co-Product	$700
LOM AISC (US$/oz Au) Net of Silver-Zinc_lead By-Products	$560
LOM AISC (US$/oz AuEq) Co-Product	$842
Capital Expenditures
Pre-Production Capital Expenditures (CDN$M)	$396
Sustaining Capital Expenditures (CDN$M)	$274
Reclamation Cost (CDN$M)	$6.5
Economics
After-Tax NPV (5.0%) (CDN$M)	$423
After-Tax NPV (7.5%) (CDN$M)	$345
After-Tax NPV (10.0%) (CDN$M)	$279
After-Tax IRR(%)	29.5
After-Tax Payback Period, discounted at 7.5% (Years)	2.7
After-Tax NPV7.5:CAPEX Ratio	0.9:1
Pre-Tax NPV (5.0%) (CDN$M)	$689
Pre-Tax NPV (7.5%) (CDN$M)	$578
Pre-Tax NPV (10.0%) (CDN$M)	$484
Pre-Tax IRR (%)	39.6
Pre-Tax NPV7.5:CAPEX Ratio	1.5:1
Average Annual After-Tax Free Cash Flow (Year 1-5) (CDN$M)	$160
LOM After-Tax Free Cash Flow (CDN$M)	$630

Cash costs are inclusive of mining costs, processing costs, site G&A, and royalties
AISC includes cash costs plus corporate G&A, sustaining capital and closure costs
Payable Gold Equivalent (AuEq) calculated by dividing net sales revenue by $1,556 (i.e., $1,561/oz Au less $5/oz Au refining costs).

NPV7.5 remains positive for changes of 25% in revenue drivers (commodity prices, grade, and recovery), capital expenditure or operating costs. After-tax economic sensitivities to commodity prices are presented in Table 1.3 illustrating the effects of varying gold price as compared to the base-case. Additional Project sensitivities will be presented in the Technical Report.

	Lower	Base	Higher
	Case	Case	Case
Gold Price (US$/oz)	$1,400	$1,561	$1,700
After-Tax NPV (5.0%) (CDN$M)	307	423	523
After-Tax NPV (7.5%) (CDN$M)	242	345	433
After-Tax NPV (10.0%) (CDN$M)	187	279	358
After-Tax IRR (%)	23.6	29.5	34.4
After-Tax Payback discounted at 7.5% (Years)	3.2	2.7	2.4
Average Annual After-Tax Free Cash Flow (Years 1-5) (CDN$M)	140	160	177

Table 1.3 After-Tax NPV and IRR Sensitivities to Commodity Prices

A program to advance the Project through a Pre-Feasibility Study and continue diamond drilling (allow 16,000 metres) for ongoing Mineral Resource expansion is recommended. The Pre-Feasibility Study should include additional metallurgy, geotechnical site assessment drilling, First Nations consultation and environmental studies at an estimated total program cost of $8,050,000. The proposed budget for the recommended 2020 program is presented in Table 1.4.

Table 1.4 Budget for Proposed PFS Program

Task Description	Cost (CDN$)
Preliminary Feasibility Study
Metallurgical Testwork	750,000
Geotechnical Mine & Site Assessment Drilling	400,000
Environmental Study Initiationand First Nations Consultation	250,000
Diamond drilling (16,000 m)	4,800,000
Pre-Feasibility Study	800,000
PFSSubtotal	7,000,000
PFSContingency at 15%	1,050,000
PFSTotal	8,050,000

2.0 INTRODUCTION AND TERMS OF REFERENCE

At the request of Mr. John Mirko, President and Chief Executive Officer of Rokmaster Resources Corp. (Rokmaster), Micon International Limited (Micon) has prepared an Updated Preliminary Economic Assessment (PEA) of the Revel Ridge Property based on a Mineral Resource Estimate prepared in January, 2020 by P&E Mining Consultants Inc. (P&E). This Technical Report discloses the results of the updated PEA.

Rokmaster is a public, TSX Venture Exchange listed junior exploration company trading on the TSX-V under the symbol "RKR" and on the OTCQB under the symbol: RKMSF. Its head office is located at 625 Howe St., Suite 1150, Vancouver, B.C. V6C 2T6; Tel: (604) 290-4647.

Mr. Richard Routledge, P.Geo., a Qualified Person under the terms of NI 43-101, conducted a site visit of the Property for the current Technical Report on June 13 and 14, 2012. A data verification sampling program was conducted as part of the on-site review.

On October 27, 2011, Micon conducted a site visit to the J&L Property for the purposes of its 2012 PEA. At that time, Micon's team attending the site visit comprised.

Christopher Jacobs, CEng, MIMMM	Project Manager/Economic Analysis
Catherine A. Dreesbach, P.E.	Senior Mining Engineer
Bogdan Damjanović, P.Eng.	Senior Metallurgist

Apart from the 2020 drilling for which results are incomplete or not yet available, there have been no material changes on the Revel Ridge Property subsequent to Micon's team visit and Mr. Routledge's June 2012 site visit, those site visits are considered to be current for the purpose of this Technical Report.

This Technical Report has an effective date of December 08, 2020.

The present Technical Report is prepared in accordance with the requirements of Canadian National Instrument 43-101 (NI 43-101) and in compliance with Form NI 43-101F1 of the Ontario Securities Commission (OSC) and the Canadian Securities Administrators (CSA).

The quality of information, conclusions and estimates contained herein is consistent with the level of effort involved in Micon and P&E's services and is based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this Technical Report. This Technical Report is intended to be used by Rokmaster subject to the terms and conditions of its contract with Micon. This contract permits Rokmaster to file this report as a Technical Report with the Canadian Securities Regulatory Authorities pursuant to NI 43-101, "Standards of Disclosure for Mineral Projects". Any other use of this Technical Report by a third party is at that party's sole risk.

2.1 SOURCES OF INFORMATION

This Technical Report is based, in part, on internal company Technical Reports, maps and technical correspondence, published government reports, press releases and public information as listed in the References Section at the conclusion of this Technical Report. Sections from reports authored by other consultants have been directly quoted or summarized in this Technical Report, and are so indicated where appropriate.

2.2 UNITS AND CURRENCY

All measurement units used in this Technical Report are metric and the currency is expressed in Canadian dollars unless stated otherwise. Gold (Au) and silver (Ag) assay values are reported in grams of metal per metric tonne (g/t Au), unless ounces per short ton (oz/st Au) are specifically stated. Location coordinates are expressed in the Universal Transverse Mercator (UTM) grid coordinates using 1983 North American Datum (NAD83) Zone 11N unless otherwise noted.

2.3 GLOSSARY AND ABBREVIATION OF TERMS

The following list shows the meaning of the abbreviations for technical terms used throughout the text of this Technical Report.

Abbreviation	Meaning

°	Degree
3-D	Three dimensional
AA	Atomic Absorption
ac	Acre
g/t Ag	Grams per tonne of silver
Ag	Silver
ARD	Acid rock drainage
AGAT	Agat Laboratories
Armex	Armex Mining Corp.
As	Arsenic
ASL	Above sea level
g/t Au	Grams per tonne of gold
Au	Gold
Au Eq	Gold equivalent
BacTech	BacTechMining Corporation
BCSC	British Columbia Securities Commission
CA	Certificate of Authorization
CDN	Canadian
CDN$	Canadian dollars
CIL	Carbon-in-leach
CIM	Canadian Institute of Mining, Metallurgy and Petroleum

Table 2.1 List of Abbreviations

Abbreviation	Meaning
cm	Centimetre(s)
Company	Rokmaster Resources Corp.
CRM	Certified reference material
CSA	Canadian Securities Administrators
Cu	Copper
Cum	Cumulative
DCF	Discounted cash flow
DDH	Diamond drill hole
DGPS	Differential Global Positioning System
E	East
EA	Environmental assessment
EIA	Environmental impact assessment
EIS	Environmental impact statement
G&A	General and Administration
g/t	Grams per tonne
GPS	Global Positioning System
ha	Hectare(s)
HMS	Heavy media separation
Huakan	Huakan International Mining Inc.
IP	Induced Polarization
IRR
ISO	Internal rate of returnInternational Organization for Standardization
Issuer
Ind.	Rokmaster Resources Corp.
k	Indicated Mineral Resources
k$	ThousandsThousands of dollars
kg	Kilograms
km	Kilometre(s)
km/h	Kilometres per hour
koz	Thousands of ounces

kt	Thousands of tonneslitre
L
LOM	Life of mine
M	Million
m	Metre(s)
$M	Millions of dollars
Ma	Millions of years
MAG	Magnetometer survey
Meas.	Measured Mineral Resources
MEM	Ministry of Mines
Merit	Merit Mining Corp.
Micon	Micon International Limited
ML/ARD	Metal leaching/acid rock drainage
mm	Millimetres
N	North
N/A	Not applicable
NAG	Non-potentially acid generating rock
NE	Northeast
NI 43-101	Canadian National Instrument 43-101
NN	Nearest Neighbour

Abbreviation	Meaning
NPV	Net Present Value
NSR	Net Smelter Return
OK	Ordinary kriging
opt	Troy ounces per ton
OSC	Ontario Securities Commission
oz/st Au	Troy ounces gold per short ton
PAC	Portable Assessment Credit
PAG	Potentially acid generating rock
Pb	Lead
PEA	Preliminary Economic Assessment Technical Report
PFS	Pre-feasibility study
Project	Revel Ridge Project
Property	Revel Ridge Property
RC	Reverse circulation drilling
QA/QC	Quality assurance/quality control
QC	Quality control
Qualified Person	Qualified Person as defined by Canadian National Instrument NI 43-101
ROM	Run-of-mine material produced during mining
S	South
Sb	Antimony
SEDAR	Website developed by the CRA, that Provides Access to Public
	Securities documents and information filed by public companies andinvestment funds in Canada
Standard	Certified reference material
t	Metric tonne(s)
t/m3	Tonnes per cubic metre
t/h	Tonnes per hour
t/d	Tonnes per day
UPEA	Updated preliminary economic assessment
WMF	Waste Management Facility
XRF	X-ray fluorescence spectrometer
Zn	Zinc

Micon and P&E have assumed that all the information and technical documents listed in the Sources of Information section of this Technical Report are accurate and complete in all material aspects. While we carefully reviewed all the available information presented to us, we cannot guarantee its accuracy and completeness. Micon and P&E reserve the right, but will not be obligated to revise this Technical Report and conclusions if additional information becomes known to us subsequent to the effective date of this Technical Report.

Although copies of the licenses and work contract were reviewed, an independent verification of land title and tenure was not performed. Neither Micon nor P&E is qualified to express an opinion on the legality of any underlying agreement(s) that may exist concerning the licenses or other agreement(s) between third parties.

The authors of this Technical report have relied largely on the documents listed in the Sources of Information and the site visit for the information in this Technical Report, however, the

conclusions and recommendations are exclusively the authors. The results and opinions outlined in this Technical Report are dependent on the aforementioned information being current, accurate and complete as of the effective date of this Technical Report and it has been assumed that no information has been withheld which would impact the conclusions or recommendations made herein.

A draft copy of this Technical Report has been reviewed for factual errors by Rokmaster. Any changes made as a result of these reviews did not involve any alteration to the conclusions made. Hence, the statement and opinions expressed in this document are given in good faith and in the belief that such statements and opinions are not false and misleading at the effective date of this Technical Report.

3.0 RELIANCE ON OTHER EXPERTS

The Authors of this Technical Report have assumed that all the information and technical documents listed in the Sources of Information section of this Technical Report are accurate and complete in all material aspects. While the Authors carefully reviewed all the available information presented to us, the Authors cannot guarantee its accuracy and completeness. The Authors reserve the right, but will not be obliged to revise this Technical Report and conclusions if additional information becomes known to us subsequent to the effective date of this Technical Report.

Although selected copies of the tenure documents, operating licenses, permits, and work contracts were reviewed, an independent verification of land title and tenure was not performed. The Authors have not reviewed or verified the legality of any underlying agreement(s) that exist concerning the claims, leases and licenses or other agreement(s) between third parties. Information on tenure and permits was obtained from Rokmaster. Selected information was verified using the BC government mining lands website https://www.mtonline.gov.bc.ca/mtov/home.do (accessed December 08, 2020).

Select technical data, as noted in the Technical Report, were provided by Rokmaster. Micon and the Authors have relied on the integrity of such data.

The authors of this Technical report have relied largely on the documents listed in the Sources of Information and the site visit for the information in this Technical Report, however, the conclusions and recommendations are exclusively the authors. The results and opinions outlined in this Technical Report are dependent on the aforementioned information being current, accurate and complete as of the effective date of this Technical Report and it has been assumed that no information has been withheld which would impact the conclusions or recommendations made herein.

A draft copy of the Technical Report has been reviewed for factual errors by the client and Micon has relied on Rokmaster's knowledge of the Property in this regard. All statements and opinions expressed in this document are given in good faith and in the belief that such statements and opinions are not false and misleading at the effective date of this Technical Report.

4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 PROPERTY LOCATION

The Property is located in the Revelstoke Mining Division in southeastern British Columbia, approximately 32 km northeast of the City of Revelstoke, BC, 420 km northeast of Vancouver, BC, and 290 km west of Calgary, AB. The Property is within the 082M-030 NTS map sheet. The location of the portal, that is located near the centre of the Property, is UTM NAD83 11N 420,719 m E, 5,681,811 m N (51° 16' 56" N and 118° 08' 12" W), see Figure 4.1.

Figure 4.1 Property Location Map

Source: GoogleEarth (2020).

4.2 HUAKAN – ROKMASTER AGREEMENT TERMS

Mineral tenure ownership is currently registered to Huakan. Rokmaster has an exclusive option to earn a 100% interest in the Property by paying Huakan an aggregate of CDN$44,200,000 in cash on the following schedule (the "Option Period"):

1. CDN$200,000 within five business days of the date on which Rokmaster has obtained TSX Venture Exchange (TSXV) acceptance of the Huakan-RKR Agreement (the "Effective Date"). Now paid – see below.
1. An additional CDN$1,000,000 within five business days of the first anniversary of the Effective Date.

1. An additional CDN$4,000,000 within five business days of the second anniversary of the Effective Date.
1. An additional CDN$6,000,000 within five business days of the third anniversary of the Effective Date.
1. An additional CDN$13,000,000 within five business days of the fourth anniversary of the Effective Date.
1. An additional CDN$20,000,000 within five business days of the fifth anniversary of the Effective Date.

In addition, to maintain the Option, Rokmaster is to complete an updated Preliminary Economic Assessment (the "Updated PEA") on the Project on or before the first anniversary of the Effective Date. If and when Rokmaster has satisfied the aforementioned Option exercise conditions, Rokmaster would have the right and option, in lieu of acquiring the Project assets, to instead acquire all of Huakan's issued and outstanding shares from Huakan's shareholders.

There are no underlying NSR Royalties on the Property.

On February 25, 2020, Rokmaster announced that it had received regulatory approval to pay, and had paid, the first CDN$200,000 option payment to Huakan.

4.3 PROPERTY DESCRIPTION

The Property is comprised of 30 mineral claims and 10 Crown Grant Lots covering a total area of 14,474.7 ha. The mineral claims cover approximately 14,310.65 ha and the Crown Grants cover an additional 164.05 ha.

Information relating to tenure was verified by means of the public information is available through the Mineral Titles Online (MTO) system at https://www.mtonline.gov.bc.ca/mtov/home. The Authors have relied upon this public information, as well as information from Rokmaster, and has not undertaken an independent verification of title and ownership of the Property claims.

The mineral claims are listed in Table 4.1 and are illustrated in Figure 4.2. The Crown Grants are listed in Table 4.2 and are illustrated in Figure 4.3.

A legal land survey of the mineral claims has not been undertaken; however, the Crown Grant Lots are legally surveyed.

The mineral claims are in good standing until between August 16, 2021 and August 1, 2025.

The annually applied tax payment due date for Crown Grants is June 30 and is payable to the BC Government. Payment is required by the due date to ensure each Crown Grant Lot is held in good standing.

Tenure	Claim	Valid UntilArea		Mining
Number	Name	Date	(ha)	Division
398402	J1	01/08/2025	25.00	Revelstoke
398403	J2	01/08/2025	25.00	Revelstoke
398404	J3	01/08/2025	25.00	Revelstoke
398405	J4	01/08/2025	25.00	Revelstoke
398406	J5	01/08/2025	25.00	Revelstoke
398407	J6	01/08/2025	25.00	Revelstoke
398408	J7	01/08/2025	25.00	Revelstoke
398409	J8	01/08/2025	25.00	Revelstoke
398410	J9	01/08/2025	225.00	Revelstoke
398411	J10	01/08/2025	300.00	Revelstoke
398412	J11	01/08/2025	25.00	Revelstoke
398413	J12	01/08/2025	25.00	Revelstoke
399179	Sage	01/08/2025	375.00	Revelstoke
399180	J13	01/07/2026	500.00	Revelstoke
399181	J14	01/07/2026	500.00	Revelstoke
399182	J15	01/08/2026	375.00	Revelstoke
401774	Brush	01/08/2025	300.00	Revelstoke
606405	Yellow Jacket	01/08/2025	161.69	Revelstoke
1070395	J&L2	09/16/2021	20.21	Revelstoke
1070401	J&L1	09/16/2021	606.17	Revelstoke
1073472	Carnes 1	12/25/2021	504.90	Revelstoke
1073473	Carnes 2	12/25/2021	545.23	Revelstoke
1073474	Carnes 3	12/25/2021	222.26	Revelstoke
1073475	Carnes 4	12/25/2021	262.84	Revelstoke
1078024	Downie 1	08/16/2021	484.46	Revelstoke
1078025	Downie 2	08/16/2021	1998.14	Revelstoke
1078026	Downie 3	08/16/2021	1938.91	Revelstoke
1078027	Downie 4	08/16/2021	725.58	Revelstoke
1078028	Downie 5	08/16/2021	1997.22	Revelstoke
1078029	Downie 6	08/16/2021	2018.13	Revelstoke

Table 4.1 Revel Ridge Mineral Claims

Note: Claim status as of December 8, 2020.

Table 4.2 Revel Ridge Crown Grant Lots

Claim Number	Claim Name	Mining Division
L 14821	Goat Fraction	Revelstoke
L 14822	Goat No. 2 Fraction	Revelstoke
L14823	Goat No. 3 Fraction	Revelstoke
L 14824	Goat No. 4 Fraction	Revelstoke
L 14825	Goat No. 5 Fraction	Revelstoke
L 14826	Goat No. 6 Fraction	Revelstoke
L 14827	View Fraction	Revelstoke
L 14828	View No.2 Fraction	Revelstoke
L 14829	Creek Fraction	Revelstoke
L7408	Aberdeen	Revelstoke

Figure 4.2 Regional Location Map

Source: Rokmaster.

Figure 4.3 Property Boundary Map

Source: Rokmaster

4.4 GENERAL REQUIREMENTS FOR MINERAL CLAIMS

To keep British Columbia mineral claims in good standing, assessment or development work is required on a claim, on or before the set expiry date. Effective July 1, 2012, all mineral claims in the province were set back to a Year 1 requirement, regardless of how many years had elapsed since their original staking. As of that date, annual work commitments were set on a four-tier schedule, as follows:

$5.00 per hectare for anniversary years 1 and 2;
$10.00 per hectare for anniversary years 3 and 4;
$15.00 per hectare for anniversary years 5 and 6; and
$20.00 per hectare for subsequent anniversary years.

Assessment work in excess of the annual requirement may be credited towards future years. Companies are permitted to pay cash in lieu of work expenditures; however, the cost is double the schedule rate above. Before their expiry, the mineral claims will require assessment work at a rate of $20 per hectare.

4.5 PERMITTING

The Revel Ridge Property is currently covered by exploration permit MX-4-500, with a $72,500 bond (placed by Huakan) in place with the Ministry of Energy, Mines and Petroleum Resources, BC, to facilitate any required reclamation. The reclamation liabilities that fall under the bond include removing the camp and workshop, covering the PAG pile with soil and seed, scarifying and seeding the campsite, portal laydown areas and access roads, and barricade the two portals.

4.6 ARMEX STATEMENT OF CLAIM

On January 17, 2018, Armex Mining Corp. (Armex) filed a statement of claim with the British Columbia Supreme Court (Vancouver Registry). Armex claims that it has a valid letter of intent with Huakan covering Huakan's J&L Property, now named the Revel Ridge Property. Rokmaster and the TSX Venture Exchange have both been informed by Armex of their statement of claim, and Huakan has notified the Company that it intends to defend the Armex action. Huakan filed a counterclaim against Armex on March 13, 2018 and Huakan has indemnified Rokmaster against any potential losses. The lawsuits have not been resolved at the date of this Technical Report.

4.7 FIRST NATIONS WITH POTENTIAL INTERESTS IN THE REVELSTOKE REGION

According to the First Nations Consultative Boundaries Map (2005), the claim areas of five First Nations overlap the Revel Ridge Property. As the map demonstrates, the Little Shuswap Indian Band, Neskonilth Indian Band, Adams Lake Indian Band, Okanagan Indian Band and the Ktunaxa Kinbasket Tribal Council assert interests in the region embracing the Revel Ridge Property. The Property is on the periphery of all five claim areas.

In 2010, the Province of British Columbia introduced a new web application to assist with the identification of First Nation claim areas. This web tool is called the Consultative Areas Database (Public), and by accessing it users can generate a list of First Nations with potential interests in lands within the province. In this instance, the Consultative Areas Database (Public) generates a report indicating that two political organizations and twelve First Nations have potential interests in the Revel Ridge Property. In the list below, the First Nations have been grouped according to their affiliations with political organizations:

Consultative Areas Database Report on the Revel Ridge Project Area.

Shuswap Nation Tribal Council (political organization not returned by CAD).

1. Shuswap Indian Band (Teit's Kinbasket band on Windermere Lake).
1. Little Shuswap Indian Band (Teit's Lake Shuswap band at Salmon Arm aka Squilax).
1. Splats'in First Nation (Spallumcheen).
1. Neskonlith Indian Band.
1. Adams Lake Indian Band.
1. Okanagan Nation Alliance.
1. Okanagan Indian Band (Northern Okanagan).
1. Penticton Indian Band (Northern Okanagan).
1. Lower Similkameen Indian Band (Northern Okanagan).
1. Ktunaxa Nation Council.
1. Akisqnuk First Nation (Upper Kutenai on Windermere Lake).
1. Lower Kootenay Band (Lower Kutenai at Creston, BC).
1. St. Mary's Indian Band (Upper Kutenai aka Fort Steele band).
1. Tobacco Plains Indian Band (Upper Kutenai at Tobacco Plains, BC).

There is an additional First Nation with a potential interest in the Property identified in the ethnographic sources: the Lakes (Sinixt) First Nation. The reason this First Nation is not returned by Consultative Areas Database as having potential interests in the Property is due to this aboriginal group being considered "extinct" by Canadian governments. There are, however, Lakes people living on the Colville Reservation in Washington State.

Exploration requiring a Notice of Work requires that the government of British Columbia to consult with all of these groups. It is the practise of Rokmaster to conduct its own First Nations consultations prior to and during its work program on the Property, as necessary. As the Project advances, more in-depth discussions and expectations should be expected. It can be expected that each group will have a different strength of claim in relation to any economic benefits discussions.

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 ACCESSIBILITY

Vehicle access to the Property is via Provincial Highway 23, north of the City of Revelstoke and Trans-Canada Highway 1. At 32 km north of Revelstoke, Highway 23 intercepts the Carnes Creek Forest Service Road. The Property is then reached by travelling eastward 13 km along the Carnes Creek Forest Service Road to the Revel Ridge Mine camp. Road travel time to the camp is approximately 45 minutes from Revelstoke. The Forest Service Road is radio controlled, however, currently is not being used for logging activities. Due to lack of activity by logging companies, road maintenance has been undertaken by Huakan and Rokmaster. Helicopter access from Revelstoke takes approximately 15 minutes. There are three helicopter bases in Revelstoke from where flights can be chartered.

On the Property, access is via four-wheel drive or tracked vehicle. The road to the camp and 832 m level portal and shop are in good condition. Several overgrown trails access the majority of the workings on the Property. The original bridges over Carnes Creek and over McKinnon Creek were destroyed by a flood in 2008. In 2008 a detour was built to a new bridge over Carnes Creek providing access to the camp. The detour starting at kilometre 10, has a locked gate controlled by the Company. The road to the original 830 m level portal is not drivable at this time due to road slumping and erosion.

5.2 CLIMATE

Revelstoke has a humid continental climate with the Koppen-Geiger classification Dfb. The average annual temperature is 6°C (https://en.climate-data.org/north-america/canada/britishcolumbia/revelstoke-714868/). The summer weather is considered moderate with July average temperatures of 18.7°C. The average annual precipitation is 103 cm/year. Winters are long and are characterized by heavy snowfalls (1 to 4 metres) with cool temperatures. Average January temperatures are -6.5°C. Snowfall typically occurs between October and May at higher elevations and between November and April at lower elevations such as at the camp and portals elevation. Exploration, development and production activities can be carried out on a yearround basis.

5.3 LOCAL RESOURCES

Revelstoke is a city with a census population of 7,547 (2016) that is located on the Trans-Canada Highway and the Canadian Pacific Railway (CP Rail). The economy of Revelstoke is forestry, construction, tourism, hydro electrical operations, transportation (mainly CP Rail) and public services. There is a large, skilled workforce of trades and technical professionals, as well as equipment suppliers available throughout the region.

The Revelstoke Hydroelectric Dam, located on the Columbia River, is 3.5 km north of Revelstoke and produces power for a large portion of British Columbia. There are no power lines running along Highway 23, although there is an underground telephone line.

5.4 INFRASTRUCTURE

Revelstoke and the surrounding area are well serviced by the Trans-Canada Highway 1 and the CP rail line. Highway 1 provides access to Calgary, located 407 km east, and Kamloops, 212 km to the west. Revelstoke has a commercial airport. The nearest airports with scheduled flights are Kelowna, BC, and Calgary, AB.

The Project assets include a rail siding and load-out facility for CP Rail in Revelstoke, a fleet of formerly utilized underground mining equipment is stored in the Company yard north of Revelstoke with a locked warehouse containing mining equipment, supplies, parts and spares that serviced the underground drifting and drilling programs of 2008–2012 and now service Rokmaster's underground drilling program.

The Property has a fully functional 40-man camp with an effective snow roof near the 832 m level portal. A water treatment plant was installed in the camp; however, it was removed in 2014 and stored in a yard north of Revelstoke. There is a large maintenance shop, dry, lunchroom, first aid and office facility, all in good condition, located in the immediate vicinity of the 832 m level portal. Electric power is produced by on-site diesel generators, enabling use of a satellite phone and internet system. A 40,000 litre Enviro-tank is currently located next to the generator shed.

The Property hosts several portals and drifts. Only two (2) portals are able to be accessed, namely the 830 m level portal and 832 m level portal. A total of 3.1 kilometres of operational underground workings are present on the Property, although access is restricted when without ventilation or pumping.

The 1,400-metre long 830 m level track drift (2.4 m x 2.4 m profile) has exposed the Main Zone for approximately 800 metres in length. This track drift was started in 1965 and has been extended on numerous occasions by subsequent owners. Huakan extended the track system in 2011/2012 by 450 m. The 830 m level track drift has not deteriorated over time and is currently in use. Five tracked crosscuts totalling 1,150 metres run northeast from the main 830 m level track drift (into the hanging wall) provided drill stations for diamond drilling that define the deposits. Raises off the 830 m level track drift have aided in the extraction of several bulk samples since the 1990s. Figure 5.1 shows a 3-D view of the underground workings, and Figure 5.2 shows a plan view of the same.

Figure 5.1 3-D View of Underground Workings

Figure 5.2 Plan View of the Underground Workings of 832 m and 830 m Level Drifts

Source: Puritch et al (2020).

The 550-metre long (5 m x 5 m profile) 832 m level trackless drift was installed by Merit in 2008 and connects to the 830 m level track drift, thereby providing year-round underground access to the 830 drift. Approximately 350 metres from the 832 portal, the decline ramps up to connect to the track drift. Due to this configuration, the 832 drift is susceptible to flooding. The 832 m level trackless drift also extends about 50 metres further as a decline from the invert. One could extend this decline a further 100 metres to drift through the Yellowjacket Zone.

Water drains from the 832 m level portal into a two-compartment settling pond outside of the portal. During the Fall of 2020, the daily average flow rate from the 832 m level portal was approximately 5 L/s. The overflow from the settling ponds gravitates about 200 m away to a flat area 350 m from Carnes Creek where it permeates into the forest floor. Water samples collected from the settling pond discharge contained an average of 0.041 mg/L total arsenic (cf an average of 0.037mg/L As between 2012 and 2019). However, during the same period, no arsenic was detected in samples collected from Carnes Creek at the old bridge crossing below the 832 Portal.

Approximately 200 metres south from the 832 m level portal is a lined Potentially Acid Generating (PAG) waste rock storage area that was constructed by Huakan in 2011 (Figure 5.3). The PAG pile is covered by tarps and was built to drain to one corner, with underflow piped into a seepage pond. Generally, the outflow pipe from the lined PAG pad is dry throughout the year but after some rains there is a trickle of outflow that allows the collection of a sample. Samples of this flow collected between 2012 to 2019 show acceptable results for BCWQ Criteria for Freshwater Aquatic Life, excepting cadmium for which total and dissolved cadmium in samples averaged 0.0003 mg/L. However, no cadmium was detected in samples from Carnes Creek collected during the same period.

Selkirk Helicopters occasionally uses the Property as a re-fuelling station for their operations. A skid mounted fuel tank is located about 300 metres from the 832 m level portal on the Forest Service Road.

There is a helicopter-accessible ski chalet located 5 km east of the Property at the lower portion of the Durrand Glacier which is used for heli-skiing in the winter and alpine hiking in the summer.

Source: Puritch et al (2020).

5.5 PHYSIOGRAPHY

The topography is characteristic of the Selkirk Mountains. The elevation ranges from 700 to 3,050 metres above mean sea level. The topographic relief is a result of recent alpine glaciation. Incised creeks, such as McKinnon Creek, created narrow valley floors, while major creeks, like Carnes Creek, exhibit a broader U-shaped appearance with the potential for deep valleybottom overburden. The talus covered slopes are steep, ranging from 28° to 40° while bedrock slopes grade up to near vertical, depending on lithology.

All of these conditions make traversing the Property somewhat hazardous and time consuming. Numerous avalanche chutes occur in the area. An avalanche chute occurred beside the original 830 portal and prompted the driving of the 832 m level trackless drift which allows safe yearround access to the underground workings. Flat ground is limited on the Property, however, there appears to be enough for a process plant site and waste rock storage piles both north and south of McKinnon Creek, should the Project advance to production. There is a tributary valley three km upstream on Carnes Creek that might be serviceable as a waste management facility, however, would require further study prior to permitting.

The main watercourse on the Property is Carnes Creek, which transects the area. Carnes Creek is around 10-25 m across and fast moving. Its main source is the Durrand Glacier, which is east of the Property. McKinnon Creek is a tributary of Carnes Creek and is a more juvenile watercourse that is between 10 to 15 m wide and can change its flow volume rapidly. These watercourses enter the Columbia River that flows southwest through Washington State to the Pacific Ocean.

The area surrounding the intersection of McKinnon and Carnes Creeks has been the focus of the majority of the work over the life of the Property and is where the camp, shop and 832 m level portal are located.

Vegetation on the Property changes from alder, devil's club, stinging nettles and deadfalls in the valley floor, through stands of cedar, hemlock and minor fir on the mountainsides, to subalpine to alpine at approximately 1,980 metres elevation. The Carnes and Tumbledown Glaciers are immediately east of the Property boundary.

6.0 HISTORY

This section is primarily based on the Property history summarized in the Technical Report by Puritch et al. (2020).

The Property area was first explored as early as 1865 when placer miners discovered gold in Carnes Creek. In 1896 prospectors, Jim Kelley and Lee George, staked the first claims at the junction of Carnes and McKinnon Creeks. The earliest work (1896-1900) carried out at the Roseberry mineral zone, 4.5 kilometres northwest of where the Main Zone was later discovered. The Property was formerly referred to as the J&L Property since its discovery by these two prospectors, Jim and Lee.

Development on the Revel Ridge Main Zone mineralization began in 1912 with the collaring of the 986 m level portal and 2 shallow shafts (each 46 metres deep). By 1924 metallurgical tests were attempting to resolve problems due to the high arsenic content of the mineralization. During 1924-27, Porcupine Goldfields Development and Finance Company completed 43 metres of underground drifting on two levels. In 1925, Mr. E. McBean excavated 30 trenches and pits along the surface trace of the Main Zone on Goat Mountain. In the following year, 26 kg of Main Zone mineralized rock were shipped to the Department of Mines in Ottawa for metallurgical testing. By 1927, the Big Bend road had reached the mouth of Carnes Creek, improving the access to the Property. The Geological Survey of Canada mapped the Property area in 1928, under the direction of Dr. H. Gunning.

Mr. T. Arnold acquired the Crown Grants and mineral claims in 1934. He, and subsequently his estate, had controlled these claims and Crown Grants until August 2010 when Merit exercised its option to own 100% interest in the Property. Between 1929 to 1933, significant development was completed on the A&E prospect, to the northwest of the Main Zone.

In 1935, Raindor Gold Mines optioned the Property and extended the 986 m level adit to 152 metres long on the Main Zone. In 1946, the two shafts were deepened, collectively to 117 metres. In 1952, Asarco optioned the Property and completed several trenches on the Main Zone. In 1962 Westairs Mines Ltd optioned the J&L, A&E and Roseberry prospects. In 1965, Westairs Ltd. collared a new portal, the 830 m level (tracked) adit to explore the Main Zone. Its total length was 297 metres. This has become one of the major underground assets on the Property. A road (12.4 km) was finally built into the Property from the Big Bend road (now Hwy 23) that same year.

In 1980, Pan American Minerals (Pan American) leased the Property from T. Arnold. In 1981, the Property was optioned by BP Minerals Ltd., Selco Division (BP-Selco) who commenced a large surface and underground exploration program. BP-Selco extended the 830 m level with an additional 1,333 metres of tracked drift and crosscuts. They completed 64 underground drill holes (2,640 metres) over the next four years. In 1986 to 1987, Noranda Mines Ltd. optioned the Property and completed metallurgical studies on the Main Zone. In 1987-88 Pan American extended the 830 m level with an additional 250 metres of tracked drift and crosscuts and completed 4 raises totalling 120 metres.

Equinox Resources Ltd. (Equinox) optioned the Property from Pan American in 1988 and completed 32 underground drill holes for a total of 2,985 metres between 1988 and 1989. A 270-ton bulk sample was mined from three TDBs (Take-Down-Backs) for metallurgical studies. Cheni Gold Mines Ltd. (Cheni) became part of the joint-venture group in 1991 with the discovery of the Yellowjacket mineralization from 32 surface drill holes. The newly discovered mineralization is situated in the hanging wall of the Main Zone and was considered a blind deposit (i.e., there is no surface evidence of the Deposit, although boulders of the Yellowjacket mineralization are present in McKinnon Creek).

In 1991, Cheni also collared a new trackless 832 m level portal (3.0 m x 3.5 m) with an adit that ran 170 metres long, stopping short of linking to the 830 m track drift. In 1991, Equinox announced a Mineral Resource Estimate (historic resource estimate) for both the Main Zone and Yellowjacket Zone. The historic Mineral Resource Estimates are not reported here since they have not been relied upon.

Metallurgical testing continued on Main Zone material through the early 1990s.

In 1996, Weymin Mining Corporation (Weymin) optioned the Property from Equinox Resources Ltd., a subsidiary of Hecla Mining Corporation. Three surface drill holes (503 metres) were completed (Table 6.1) and a 120-tonne underground bulk sample was retrieved from the 830 m level for metallurgical studies from six sample locations.

Drill Hole	From	To	Core Width	Au	Ag	Pb	Zn
No.	(m)	(m)	(m)	(g/t)	(g/t)	(%)	(%)
S-97-1	92.42	95.28	2.86	0.09	18.08	0.89	3.22
	95.28	97.28	2.00	0.00	5.66	0.29	0.08
	98.28	99.28	1.00	0.00	40.46	0.83	0.95
S-97-2	67.15	68.22	1.07	0.07	64.46	3.05	11.94
	75.72	80.50	4.78	0.24	63.06	2.38	14.92
	84.05	86.05	2.00	0.15	33.09	1.98	5.68
S-97-3	75.06	75.54	0.48	0.00	27.77	1.63	3.80
	82.54	87.02	4.48	0.09	52.71	2.43	11.10
	93.55	94.38	0.83	0.34	121.78	5.69	15.19
	96.07	98.92	2.85	0.23	29.10	1.58	5.87

Table 6.1 1996 Yellowjacket Zone Drill Highlights

In 1996, Weymin commissioned H.A. Simons of Vancouver to complete two detailed reports: "Technical Review of the J&L Property" and "Project Opportunities for the J&L Property". In March 1998, H. A. Simons completed the "McKinnon Creek Property Scoping Study". Simons provided analyses of six cases, exclusively on the Main Zone. The Yellowjacket Zone was not analyzed. The two favoured cases are not reported here since they are historical and are not relied upon.

BacTech Mining Corporation (BacTech) optioned the Property in 2004. BacTech carried out further metallurgical tests, engineering and environmental studies. A minor drilling program

was carried out that year. Due to the financial collapse of BacTech, the drilling details have never been made available.

On April 13, 1997, Merit entered into an option agreement with the Estate of T. Arnold to acquire a 100% interest in the Property. By December, 2007, a 40-man camp was installed, construction of a shop/mine dry complex was completed, and mining equipment was procured. A late fall 2007 surface diamond drilling program of nine holes, totalling 1,363.37 metres, was completed, with the objective of verifying historic drilling over a portion of the Yellowjacket deposit. The program successfully achieved this objective by intercepting multiple zinc-leadsilver zones similar in grade and width to previous drilling, summarized in Table 6.2.

Drill	From	To	Length	Ag	Pb	Zn	Combined
Hole No.	(m)	(m)	(m)	(g/t)	(%)	(%)	Pb-Zn (%)
M07SJ-01	22.00	25.00	3.00	11.80	0.65	2.83	3.48
	27.50	30.00	2.50	25.92	1.32	6.83	8.15
	33.10	34.85	1.75	84.55	3.01	12.07	15.08
	39.35	41.15	1.95	61.00	3.49	3.39	6.88
	43.90	63.30	19.40	27.85	1.10	5.19	6.29
Including	43.90	49.50	5.60	27.06	1.12	8.54	9.66
	Incl. 50.90	57.55	6.65	22.01	0.93	4.40	5.33
	Incl. 58.80	63.30	4.50	52.19	1.89	5.06	6.95
	67.50	69.00	1.50	58.13	3.63	9.65	13.28
M07SJ-02	23.15	28.00	4.85	44.58	1.75	7.22	8.97
	31.30	36.10	4.80	47.31	1.97	4.77	6.74
	40.90	43.00	2.10	22.72	1.13	3.91	5.04
	50.15	58.00	7.85	9.98	0.38	4.11	4.49
	68.60	73.00	4.40	55.40	2.47	9.65	12.12
	98.00	99.00	1.00	66.20	1.64	9.54	11.18
M07SJ-03	34.00	37.00	3.00	55.57	1.91	8.43	10.34
	46.00	46.75	0.75	108.00	4.33	7.21	11.54
	52.00	54.70	2.70	65.08	1.89	12.71	14.60
	61.05	61.55	0.50	66.20	1.64	9.54	11.18
	68.00	70.00	2.00	19.15	1.04	10.75	11.79
	72.00	73.50	1.50	30.53	1.80	8.01	9.81
M07SJ-04	48.30	49.65	1.35	104.36	4.97	7.75	12.72
	109.75	111.15	1.40	0.80	0.03	4.83	4.86
	116.00	120.00	4.00	156.88	0.56	1.09	1.65
M07SJ-05	40.60	43.00	2.40	38.82	1.57	4.50	6.07
	45.00	49.05	4.05	11.57	0.64	2.93	3.57
	52.00	55.20	3.20	44.94	1.98	20.05	22.03
	58.00	59.50	1.50	102.73	2.83	17.93	20.76
	98.00	99.35	1.35	38.30	1.23	5.78	7.01
M07SJ-06	31.00	47.00	15.52	56.08	2.28	6.11	8.39
Including	31.00	36.00	5.00	74.14	2.70	8.66	11.36
Including	38.90	47.00	8.10	62.85	2.73	6.61	9.34
	61.25	63.00	1.75	66.55	2.72	10.14	12.86
M07SJ-08	95.75	98.40	2.65	41.30	1.69	4.14	5.83

Table 6.2 2007 Yellowjacket Zone Drill Highlights

DrillHole No.	From(m)	To(m)	Length(m)	Ag(g/t)	Pb(%)	Zn(%)	CombinedPb-Zn (%)
M07SJ-09	91.00	92.30	1.30	113.23	0.77	3.64	4.41
	100.20	101.50	1.30	13.40	0.87	5.08	5.95

The 2007 surface drilling program also intercepted Main Zone material; however, the Main Zone is not strongly developed adjacent to the Yellowjacket north area and ranges from 0.25 to 3.75 metres wide with lower metal values.

Rehabilitation of the 832 m portal and underground development commenced in January 2008. The original 170-metre long Cheni 832 drift was slashed out to a 5 m by 5 m profile to allow for the passage of 30 tonne trucks. The 832 m level drift was extended a further 550 metres with the 5 m by 5 m profile, and connected to the 830 m level track drift approximately 310 metres from the original 830 m level portal. This allowed for easy underground access. This drifting was completed by September 2008, at which time the program was suspended, due to financial constraints and a major downturn in world metal prices.

Resumption of mineral exploration activity at the Property by Huakan began in November 2010, with the implementation of the 2010-2011 winter underground drill program aimed at verifying historic drilling and generating a NI 43-101 Mineral Resource Estimate.

Between November 15, 2010 and January 30, 2011, Huakan completed 60 underground diamond drill holes for a total of 7,873.74 metres of BQTW core. The program started as a 12 hole in-fill drill program but was extended to expand the edges of the Main Zone deposit at 30 metre centres. By May 16, 2011 Huakan announced an NI 43-101 Mineral Resource Estimate on the Main Zone with a Technical report prepared by P&E Mining Consultants Inc., filed June 23, 2011.

Huakan subsequently engaged Micon to prepare a PEA report, utilizing the May 16, 2011 Mineral Resource Estimate. The results of the PEA were announced on April 24, 2012, with the PEA report filed on SEDAR on June 6, 2012.

In 2012 Huakan conducted a 450-metre drifting and a 45-hole, 9,725 metre underground drill program to expand the Mineral Resource Estimate of the Main Zone. The 2012 program was successful in increasing the Mineral Resources and results of an Updated Mineral Resource Estimate by P&E were reported in a news release by Huakan dated September 18, 2012. This estimate significantly increased Indicated Mineral Resources on the Main Zone and for the first time included a Mineral Resource Estimate on the Yellowjacket Zone. No subsequent physical work has been done on the Property since the 2012 Updated Mineral Resource Estimate. In January, 2013, Huakan reported updated metallurgical test work results from a bulk sample collected in the 2012 program.

All Huakan 2012 drilling was done with wireline BQTW diamond core. True widths are approximately 75% of downhole intercept lengths. The mineralization dips NE at approximate 56o . Core recovery was >90% and often >95%.

From the summer of 2010 until the spring of 2014, Huakan conducted an extensive campaign of metallurgical testwork on bulk sample material from Main Zone and some testwork on Yellowjacket Zone material, from core. Test work included comminution testing, heavy media separation, open cycle flotation on Main Zone with optimization and variability testing and open cycle flotation on Yellowjacket Zone, Lock cycle flotation on both Main Zone and Yellowjacket Zone, flotation tailing characterization, and treatment and cyanidation of gold concentrate by bioleaching and pressure oxidation.

Commencing in May, 2020, Rokmaster undertook road maintenance, permitting, and rehabilitation underground and on surface in preparation for underground drilling. Underground drilling commenced September 23, 2020 and continues to date. The results of that work were not available for inclusion in this PEA, however.

6.1 PREVIOUS MINERAL RESOURCE ESTIMATES

A previous Mineral Resource Estimate for the Property with an effective date of May 16, 2011 was reported at an NSR cut-off grade of CDN$110/tonne (Table 6.3).

Classification	Tonnes	Au(g/t)	Au(ozs)	Ag(g/t)	Ag(ozs)	Pb(%)	Zn(%)
Main Zone
Measured	1,202,000	6.71	259,200	69	2,664,600	2.4	4.46
Indicated	1,165,700	6.92	259,200	64.9	2,432,100	2.01	3.86
Measured & Indicated	2,367,700	6.81	518,400	66.95	5,096,700	2.21	4.16
Inferred	4,538,100	5.19	757,500	67.8	9,887,800	2.16	2.99
Footwall Zone
Inferred	292,800	4.54	42,700	49	461,900	0.91	0.73

Table 6.3 Previous 2011 Mineral Resource Estimate

Source: Brown, F., Ewert, W., Armstrong, T. (2011), Technical Report and Resource Estimate J&L Property, Revelstoke BC Canada. Technical report prepared for Huakan International Mining Inc., with an effective date of May 16, 2011.

A Qualified Person has not done sufficient work to classify the above historical estimate as a current Mineral Resource. The Issuer is not treating the historic estimate as a current Mineral Resource and it should not be relied upon.

The May 16, 2011 Mineral Resource Estimate was updated and published by a press release on September 18, 2012 to include the results of the 2012 drilling program (Table 6.4).

Classification	Tonnes	Au(g/t)	Au(ozs)	Ag(g/t)	Ag(ozs)	Pb(%)	Zn(%)
Main Zone
Measured	1,313,000	6.37	268,800	65.1	2,747,000	2.26	4.22
Indicated	2,640,000	5.34	453,200	52.2	4,432,000	1.78	3.23
Measured & Indicated	3,953,000	5.68	7,222,000	56.5	7,179,000	1.94	3.56
Inferred	4,337	4.16	580,200	57.8	8,057,000	1.82	2.72
Footwall Zone
Inferred	363,000	3.65	42,500	25.4	296,000	0.55	0.51
Yellowjacket Zone
Indicated	1,003,000	0.21	6,900	64.1	2,068,000	2.77	9.08
Inferred	35,000	0.35	400	81.9	91,000	3.18	6.26

Table 6.4 Previous 2012 Mineral Resource Estimate

The September 18, 2012 Mineral Resource Estimate was updated and published by a press release on January 23, 2018 (Table 6.5).

TotalAll Zones	Tonnes(k)	Au(g/t)	Au(koz)	Ag(g/t)	Ag(koz)	Pb(%)	Zn(%)	AuEq(g/t)	AuEq(koz)
Measured	1,337	6.19	266	63.3	2,721	2.21	4.12	9.69	417
Indicated	3,823	4.03	495	53.0	6,509	1.98	4.73	7.60	934
Meas & Ind	5,160	4.59	761	55.6	9,231	2.04	4.57	8.14	1,351
Inferred	4,808	4.35	672	60.6	9,367	1.84	2.55	6.95	1,075

Table 6.5 Previous 2018 Mineral Resource Estimate

Note: k = thousands, koz = thousands of ounces.

All historical Mineral Resource Estimates have been superseded by the updated Mineral Resource Estimate that is the subject of this Technical Report. This Technical Report updates the previous Mineral Resource Estimates by incorporating changes in the commodity prices. No additional drilling or sampling information was used.

The current 2020 Mineral Resource Estimate can be found in Section 14.0 of this Technical Report.

7.0 GEOLOGICAL SETTING AND MINERALIZATION

This section is primarily based on the Property geology summarized in the Technical Report by Puritch et al. (2020).

7.1 REGIONAL GEOLOGY

The Property lies within the Selkirk Mountains near the north end of the Kootenay Arc, that is a complex sequence of northwest trending, east dipping Neoproterozoic to Lower Paleozoic metasedimentary and metavolcanic rocks of the North American craton (Logan et. al., 1996, 7 A & B). The belt is characterized by tight to isoclinal folds and generally west verging thrust faults. Greenschist grade regional metamorphism has affected most of the rocks in the map area. Recent mapping by provincial government geologists has outlined the regional geology of the area.

7.2 PROPERTY GEOLOGY

The Property is underlain by north to northwest striking, moderate to steeply east-dipping metasediments and metavolcanic rocks of the Neoproterozic/Lower Cambrian Hamill Group, overlying Lower Cambrian rocks of the Mohican and Badshot Formations, and Lower Paleozoic Lardeau Group rocks. These units consist, for the most part, of sheared to intensely folded impure quartzites, quartz sericite to sericite to chlorite schists and phyllites, and grey banded to carbonaceous limestones.

The following is a brief description of the main geological units that are present on the Property. A stratigraphic column displaying the age relationships of units is presented in Figure 7.1 and a map of regional geology in Figure 7.2.

7.2.1 Hamill Group

The Hamill Group rocks are predominantly interbedded medium brown to green-black sericitic and/or chloritic quartzites and phyllites with minor layers of argillite and graphite. This unit appears as the upper Hamill unit described by Logan et.al., 1996, and is probably Lower Cambrian in age. Hamill group rocks form part of the footwall and hanging wall of the Main Zone deposit. The unit has a gradational upper contact with the Mohican and Badshot Formations.

7.2.2 Mohican Formation

The Mohican Formation is Lower Cambrian in age (Fritz et. al., 1991). This unit is located at the eastern and southern boundary of the original J&L claims. The eastern unit is in the hanging wall of the Main Zone. It is characterized as limonite-rich, sericitic chloritic calcareous phyllite and quartzite interlayered with narrow layers of marble. Logan et al. (1997A) describes the Mohican Formation as a "transition between quartz-rich sediments of the Hamill Group and the carbonate-rich rocks of the Badshot Formation".

Figure 7.1 Stratigraphic Column

Source: Rokmaster

Figure 7.2 Revel Ridge Regional Geology Map

Source: Rokmaster.

Figure 7.3 Revel Ridge Site Geology Map

Source: Rokmaster.

7.2.3 Badshot Formation

The Badshot Formation is the most visible and distinctive lithologic unit within the claims and is Lower Cambrian in age. This white to grey, fine to medium-grained limestone/dolomite/marble and varies in its silica content. The Yellowjacket Zone is contained within this Formation. The higher silica content of the Yellowjacket appears to be alteration specific to the Yellowjacket mineralizing system. The Main Zone crosscuts the Badshot Formation as observed in the 830 m level tracked drift. Several diamond drill holes display good grades and widths where the Main Zone crosscuts the Badshot Formation. Thin interlayers of black graphite are seen within the Badshot at the 832 m level portal.

7.3 LARDEAU GROUP

7.3.1 Index Formation

The Index Formation can be subdivided into at least four units (i.e., black phyllite, marble, greenstone and quartz breccia), however only two units have been identified on the Property.

The black phyllite unit is in the footwall of the Main Zone. Logan has also traced the unit in the northern portions of the claims around the A and E showings. The unit can be calcareous and graphitic and may contain minor marble and quartzite layers.

The greenstone unit within the Property is observed as a series of diorite sills. The diorite is composed predominantly of coarse-grained chlorite and plagioclase feldspar. The closest sill is approximately 600 metres northwest of the North Zone Pit (approximately 500 metres northwest of the intersection of the Main Zone with McKinnon Creek). Another diorite sill is immediately east of the Roseberry showing. A third sill is at the summit of Goat Mountain.

7.3.2 Micaceous Quartzite Unit

The Micaceous Quartzite Unit is predominantly at the western edge of the Property and is well exposed along the Carnes Creek Forest Service Road. The unit is composed predominantly of quartzites to siliceous phyllites to quartz muscovite schists and may be loosely correlated to the Broadview Formation (Brown, 1991).

7.3.3 Jowett Formation

This Jowett Formation is exposed in the first kilometre of the Carnes Creek Forest Service Road. It is an interlayered, green, metavolcanic and non-carbonaceous marble. This unit forms the hanging wall of the Columbia River Fault in the area of the claims.

7.4 LOCAL GEOLOGY

Proximal to the Main Zone, the lithological assemblage consists of phyllite and schist (87%), limestone (8%), quartzite (5%), and rare dykes as defined by core logging from the 2010/2011 and 2012 drill campaigns.

The phyllite and schist units are moderately to well foliated, consisting of variable amounts of sericite, chlorite, and quartz. Chlorite, though in minor amount, is considered the major contributor of the distinctive green hue in the units. Some banded sericite-chlorite-phyllite zones, ranging in width from 0.5 to 2.0 metres, have a distinctive brownish hue due to the presence of fine-grained biotite. Although the phyllite is highly sheared and strongly foliated, feldspar phenocrysts are noted in the core indicating a possible mixing of a volcanic and/or sedimentary protolith.

There are two types of limestone seen in core proximal to the Main Zone, namely carbonaceous limestone and banded limestone, varying in bands with widths from 1.0 to 20.0 metres. The carbonaceous limestone units are fine to medium grained, dark grey to black in color, weakly to moderately foliated, and intensely jointed. The banded limestone units are a light grey and medium grey, medium-grained, moderately to well inter-banded, recrystallized limestone.

Quartzites are clean generally milky white in color, fine-grained and massive to weakly banded, with minor sericite and/or chlorite on foliation planes.

Rare dykes are present as late-stage porphyritic intrusions. They are dark greyish green to brown in colour and medium-grained, composed of feldspar, quartz, and varying amounts of biotite. Only one dyke occurrence is observed in one drill hole from the 2010/2011 drill program. Its upper and lower contacts were sharp.

7.5 STRUCTURE

The dominant orientation of structure fabric and lithologic contacts on the Property strikes northwesterly (striking about 330°) and dips (about 50°) toward the northeast. Near the southern edge of the Crown Grants the lithology changes strike to a more east-west orientation (striking about 290°) and dipping northeast (about 40°). This change may be part of the Carnes Creek anticline (Logan et. al., 1997A) or late-stage deformation.

The rocks in the area are faulted and intensely folded. One penetrative foliation is developed in all rock types and is the most readily recognizable feature. An earlier stage foliation is observed in silicified phyllite or quartz schist. Early-stage deformation features are rarely preserved due to intense folding and strong shearing.

The Badshot Formation in the vicinity of the known deposits is recumbently overturned (Logan et. al., 1996, 1997A). This unit evidently flowed during deformation and formed boudinage structures that cannot be easily correlated with each other. Some folding structures can be seen underground but is and usually confined to the Main Zone wall rocks, not affecting the Main

Zone mineralization. Limestone is strongly folded, whereas argillite, quartzite and mineralization only locally exhibit folds where completely enclosed by folded limestone.

Two thrust sheets have been identified in the Project area striking northwesterly and dipping east (Logan et. al., 1997A). The Main Zone is also a reverse fault and is approximately strike parallel to the larger thrust faults noted to the southwest and northeast of the Main Zone. Very minor slip surfaces may sometimes be observed along the footwall and hanging wall contacts of the Main Zone. Locally these faults splay into either wall, carrying the mineralized zone with them. Displacement along the faults is generally minor.

The Main Zone is hosted in a large planer deformation zone which most often demonstrates reverse and sinistral kinematics. The deformation zone hosts the auriferous sulphides which form the Main Zone, displaying exceptional strike and down dip continuity. The deformation zone crosscuts lithologic boundaries at low angles along strike and down dip. The shear zone is preferentially developed near the contact between the limestone and phyllite or between quartz-rich schist and phyllite. Limestone tends to occur on the footwall of the mineralized zone along about half of the exposed underground strike length.

For much of the Main Zone exposed along strike underground, the zone is quite tabular with parallel sheeted massive and stringer sulphide bands. There are segments along strike where the banded massive sulphide units within the zone exhibit complex deformation textures. There are a number of indicators of shear sense, such as stretching lineation, rotated clasts, sheath folds, and asymmetric micro-folds. S/C fabrics, clast rotation and asymmetrical folds generally suggest sinistral and reverse movement senses.

The silver-lead-zinc-rich Yellowjacket Zone is considered to be a structurally controlled carbonate replacement deposit composed of multiple parallel siliceous sphalerite-galenabearing zones preferentially located at the contact between alternating units of quartzites, phyllites, and limestone. Currently, the Yellowjacket Zone has not been shown to be as laterally extensive as the Main Zone but, pending further exploration, it remains open to the northwest and down plunge. The Yellowjacket Zone sub parallels and is in the immediate hanging wall of the Main Zone.

7.6 MINERALIZATION

7.6.1 Main Zone

The Main Zone is a structurally controlled orogenic gold deposit characterized by precious metals enhanced sulphides containing significant gold, silver, lead and zinc. The deposit has a very predictable geometry. The zone is sheet-like or tabular with an average dip of about 56° to the northeast. The zone of sheeted massive and stringer sulphides having an average true width of 2.5 metres but can reach 15 metres in true thickness. The continuity of the zone is interrupted in a few places where it pinches out completely within narrow stretches. Exploration has confirmed persistent vertical and horizontal continuity of the Main Zone. On surface, a strike trace of at least 3.3 km has been defined. In addition, it is speculated that the

Main Zone is linked to the Roseberry Prospect and also to the former Mastodon Mine, which would suggest a collective potential strike length of nine kilometres. The deposit has been traced 1,500 metres along strike and 800 metres down dip by underground drilling. Within this, it is exposed for 850 metres along strike by underground drifting along the 830 Level. Extensive drilling has indicated a traceable continuous plane with no significant fault offsets, cut-offs or fault drags zones, and there is possibly an element of improved grade in en-echelon series of northwest plunging lenses that strengthen with depth.

There remains excellent potential for delineation of additional Mineral Resources on the Main Zone that remains open up and down dip, and along strike to the northwest and southeast.

The Main Zone is composed of closely spaced bands of massive sulphides that frequently coalesce at its widest parts. Individual bands, that are generally tabular, may die out along strike over tens of metres but appear to resume in an adjacent band. Individual massive sulphide bands frequently range from five centimetres to one metre thick. Sulphide minerals include pyrite, pyrrhotite, gold-bearing arsenopyrite, iron-rich sphalerite (blackjack), galena, tetrahedrite and trace chalcopyrite. There are also traces of silver-lead-antimony (Ag-Pb-Sb) and lead-antimony sulphosalts. The banding ranges from predominantly arsenopyrite (high gold), to mixed arsenopyrite and massive sulphides, to massive sphalerite with no arsenic present. Where the mineralization narrows, it is almost completely composed of arsenopyrite. Mineralization widens and sulphide assemblage is more diverse where it is in contact with, or completely enclosed by, limestone. Between mineralized bands, the host rock has been altered (sericite-quartz) and contains disseminated mineralization or thin massive to stringer sulphide streaks.

Three distinct types of mineralization have been noted:

Type I: mineralization is comprised of massive bands, lenses and stringers of sulphides in a sericitic shear zone. Sulphides consist of medium to coarse grained pyrite, variously grain sized arsenopyrite, and fine-grained fracture-filled sphalerite and galena. Some coarse-grained pyrite and arsenopyrite display a brecciated texture.
Type II: mineralization is characterized by "milled" massive sulphide texture consisting of fine to coarse-grained, rounded to sub-rounded pyrite, arsenopyrite, quartz, and wall rock clasts in a very fine-grained sulphide matrix. The matrix is composed of fine-grained pyrite, arsenopyrite, sphalerite, galena and quartz. Clasts derived from the host rock such as phyllite and schist contain sulphide stringers, which in part may represent Type I form of mineralization. This milled feature is interpreted as a mylonite texture developed within a shear zone. Milled sulphides carry high values of gold, silver, lead and zinc, and elevated mercury and antimony.
Type III: mineralization consists of narrow stringers and fine to medium-grained disseminations of principally sphalerite, with lesser amounts of galena and pyrite and very little arsenopyrite. Sphalerite is red to honey yellow in color and appears to replace limestone. Although Type III mineralization can reach widths of 6-10 metres, it appears to have limited extent both along the strike and vertically.

The Main Zone is sericite, carbonaceous and chlorite altered. The chloritic and sericitic phyllites of the wallrock are gradational in composition laterally and vertically making subdivision difficult.

Principal gangue minerals to the Main Zone include quartz, calcite, siderite, sericite and chlorite.

7.6.2 Hanging Wall and Footwall Zones

The wall rock in the hanging wall and footwall is mostly composed of sericite, chlorite, phyllite, quartz, schist, and limestone. Phyllite and schist contain 1-5% pyrrhotite in the form of micro lenses on the foliation. An increase in pyrite development, concurrent with a sharp decrease in pyrrhotite, occurs in close proximity to the mineralized zone. Phyllite and schist are bleached due to sericitic alteration and silicification, resulting in apparent colour contrast between altered and unaltered rocks. Pervasive sericitization is extensively developed within the shear zone and its immediate hanging wall and footwall. The sericitic selvage ranges from 2 to 30 metres wide. Marblization occurs immediately at the contact between limestone and the margins of the mineralized zone, varying in width from 0.1 to 1 metre.

Pyrrhotite is disseminated ubiquitously throughout much of the non-mineralized rock in minor amounts. Trace amounts of chalcopyrite and pyrite are observed.

Sub parallel intermittent footwall and hanging wall zones, similar to previously described mineralization Types I, II and III, occur proximal to the Main Zone. One hanging wall zone (HM1) (named HW Zone in the Mineral Resource Estimate) lies approximately five metres to the hanging wall of the Main Zone and has a degree of continuity in order that a Mineral Resource can be estimated. A footwall zone (FM1) (named FW Zone in the Mineral Resource Estimate) lies approximately five metres to the footwall of the Main Zone and also has sufficient continuity for a Mineral Resource to be estimated. Other zones include a second hanging wall zone (HM2) that lies approximately 20 metres to the hanging wall of the Main Zone, and a second footwall zone (FM2) that lies approximately 20 metres to the footwall of the Main Zone. Mineral Resources have not been defined on these secondary zones.

7.6.3 Yellowjacket Zone

The Yellowjacket Zone does not outcrop and as such was only discovered in 1991, late in the exploration history of the evaluation of the Main Zone. The Yellowjacket Zone is thought to be a stratabound carbonate hosted, lead-zinc-silver deposit but careful examination of the 2012 drill core of the Yellowjacket strongly supports a structurally controlled replacement model. Mineralization occurs at contacts between limestone and meta-volcanics and, because there are frequent alternations or interbedding of these two lithologies in this area, the Yellowjacket Zone is composed of multiple subzones.

The Yellowjacket Zone is generally sub-parallel to the Main Zone and is located approximately 30 metres into the hanging wall rock of the Main Zone. The lead-zinc-silver mineralization is

confined to multiple discrete zones hosted in siliceous carbonate units. The Yellowjacket Zone does not outcrop and is defined only by drilling. Limited drilling (42 holes) has traced the deposit along strike for 500 metres where the Yellowjacket Zone Mineral Resource is defined. The deposit appears to rake to the southeast at 30° (see the surface trace of these zones in Figure 4.2). The Yellowjacket Zone remains open beyond the limits of the Mineral Resource, both laterally to the northwest and at depth. Three drill holes from one collar set-up 300 metres northwest, picked up Yellowjacket Zone mineralization (Table 6.1). A 2014 soil geochemical survey indicated elevated Zn-Pb-Ag values along trend of the zone for 800 metres further northwest of the Mineral Resource area (see Figure 4.2).

The Yellowjacket Zone has no arsenic content and little gold. The mineralization is composed of disseminated and patchy massive zinc-rich honey-coloured (yellowjack) and red coloured sphalerite with minor medium-grained disseminated and hairline stringers of galena with elevated silver values. Other minerals include calcite, silica and minor sericite and siderite. Texturally, the mineralization can be foliated and/or laminated with sphalerite and galena running along cleavage surfaces. Other textures include brecciated or lacework patterns. Dolomite sections show discontinuous banding and are usually lower in grade.

The carbonate units hosting the Yellowjacket Zone may be occurring in the hinge of a recumbent isoclinal fold, fringed by phyllite and quartzite. The mineralization appears to thicken in the apparent fold hinge where darker coloured sphalerite and coarser and more abundant galena occurs. The Yellowjacket Zone is intensely silicified. Sericite has also been observed in core samples. Silicification also appears to intensify towards the apparent fold hinge. Fluorite is common in most mineralized sections, particularly near higher grade sections. Pyrite and pyrrhotite are present in low amounts.

7.6.4 Other Showings

The Roseberry showing lies on the Property (MinFile number 082M 091), 4.5 kilometres to the northwest of the Main Zone. The precious and polymetallic (Cu-Zn-Pb-Ag-Au) vein-type showing lies just below the contact of Lardeau graphitic schists and Badshot Formation limestones. The showing's existence has been known for over a century, however, it has received only minor surface exploration due to its remote location. The mineralization is composed of coarse disseminated to semi-massive arsenopyrite in discontinuous quartz carbonate veins hosted by intensely sheared graphitic schist. The mineralization resembles the Main Zone mineralization. Chip sampling of the Roseberry Showing returned values such as 15.03 g/t Au and 37.4 g/t Ag across 0.3 metres.

The A&E showing lies on the Property (MinFile number 082M 099), five km north of the Main Zone and two km northeast of the Roseberry Showing. Mineralization is related to sheared schistose zones with intense deformation and complex folding, interlayered with or in contact with limestone. This precious and polymetallic (Ag-Pb-Zn-As) showing represents a series of three parallel mineralized zones that appear similar to the Main Zone. One of the zones averaged 11.01 g/t Au, 356.7 g/t Ag, 10.75% Zn and 5.48% Pb from four muck samples. It is a narrow arsenical zone of massive sulphides. There are several hand-tooled short adits and

surface showings that have traced the zone for 400 metres along strike and 160 metres vertically. The A&E showing lies on a different horizon than the Main Zone and as such indicates potential for parallel zones of mineralization at multiple horizons.

The Copper Zone is located 100 metres to 150 metres into the footwall of the Main Zone. It is a narrow stringer sulphide zone hosted by quartzites and chloritic phyllites and schists, and has been traced for 320 metres horizontally and 90 metres vertically. Although it does not appear to return economic grades at surface, the showings are leached and weathered. A chip sample taken by Equinox returned 3.55 g/t Au, 21.7 g/t Ag and 0.19% Cu over 1.0 metre. The Copper Zone could be tested further by diamond drilling.

Figure 7.4 presents a plan showing the surface traces of mineralized zones on the property.

Figure 7.4 Plan showing Surface Traces of Mineralized Zones

Source: Rokmaster.

8.0 DEPOSIT TYPES

The Revel Ridge Property lies at the northern end of the Kootenay Arc which is known for its Irish-type carbonate hosted Zn-Pb, VMS (Gold stream) and Sedex deposits. The two main deposits known on the Property are the Main Zone and the Yellowjacket Zone.

BC Minfile No 082M 003(2012) reports that intense deformation of the deposit has distorted or destroyed most original mineralized textures and mineralized-wallrock relationships. Most textures now observed result from an overprinted tectonic fabric, making interpretation of the timing and environment of deposition difficult, at best. There are two schools of thought on the deposit classification. Early interpretations classed the deposit as an epigenetic shear zone replacement, or vein deposit. Other proponents support a syngenetic sedimentary-exhalative origin. The deposit exhibits characteristics of both models and the debate is not resolved.

8.1 MAIN ZONE

Main Zone mineralization crosscuts lithologies at low angles and follows a continuous, planar thrust fault or deformation zone. The Zone is continuous and tabular. Surface exploration has traced the Main Zone for at least three kilometres. The drilling to date demonstrates that mineralizaton in the Main Zone continues for a strike length of at least 1,500 metres and in the down dip direction for at least 800 metres. There remains good potential for additional Mineral Resources on the Main Zone, which remains open in the down dip direction and along strike to the northwest and possibly to the southeast. The Main Zone averages 2.5 metres true thickness of sheeted sulphide veining but the sheeting can reach up to 15 metres true thickness. It is a complex banded zinc-lead-silver-gold-arsenic-pyrite deposit, partly having a close spatial relationship with limestone. Massive to semi-massive bands and stringers parallel or subparallel the dominant shear foliation, reflecting the strong structural influence.

Previous workers believed that the Main Zone formed at the contact of footwall limestone with hanging wall phyllites. The 2010/2011 and 2012 drill programs suggested that the Main Zone is most likely to occur at the base of the limestone unit in contact with footwall phyllites.

The close spatial relationship between mineralization and limestone can be interpreted to be due to four factors. Firstly, the contact between limestone and phyllite may be favourable for the development of a shear zone. Secondly the competency contrast between limestone and phyllite creates a favorable condition to allow dilation within a shear zone. Thirdly, limestone is a good hydraulic seal. Fourthly, carbonaceous limestone is a good reducing agent which may be chemically favourable to the process of gold precipitation.

Based on the detailed geological core logging, the mineralization is stronger when the shear zone cuts the phyllite unit rather than more schistose lithologies.

The Main Zone does not easily fit into a specific genetic model. Geologists who have worked on the Main Zone in the past have proposed a Sedimentary Exhalitite (Sedex) model, a Volcanogenic Massive Sulphide model (VMS), a replacement model and a shear hosted model.

The Main Zone mineralization lies in an inferred shear zone, characterized by sheeted massive to semi-massive sulphide bands and stringers, spatially associated with the Yellowjacket Zone. Huakan geologists interpreted the Main Zone to be a shear hosted sheeted sulphide replacement deposit within a thrust zone, post-dating the Yellowjacket Zone mineralizing episode.

8.2 YELLOWJACKET ZONE

The Yellowjacket Zone is interpreted alternatively as a carbonate replacement deposit (CRD) or Mississippi Valley Type (MVT) deposit with significant metal values in Ag, Zn and Pb.

Most CRD or MVT Ag-Pb-Zn deposits within the Kootenay Arc, including the HB, Jersey-Emerald, Reeves-Macdonald, and Mastodon, are a few million tonnes in size with the largest of these in the 6–10 million tonne range and contain and average of 3-4% Pb, 1-2% Zn, and 2–4 ppm Ag, All Kootenay Arc deposits are hosted in the Cambrian platformal carbonate Badshot Limestone. These deposits are distinctly epigenetic and (based on Re-Os dating) formed during the Devonian, significantly post-dating the lower Cambrian age of the enclosing stratigraphy (Paradis and Simandl, 2010).

In contrast to most of the Kootenay Arc Pb Zn deposits, however, the Yellowjacket Zone exhibits much higher silver enrichment and significantly higher overall Pb-Zn values with the indicated mineral resource currently standing at 771,000 tonnes averaging 0.09 g/t Au, 62.6 g/t Ag, 2.6% Pb and 9.93% Zn (Puritch et al., 2020). Moreover, the Yellowjacket is hosted by carbonate rocks which likely belong to the Hamill Group, an older stratigraphic package forming the footwall to the Badshot Limestone.

Unlike the Main Zone, gold and arsenopyrite contents within the Yellowjacket Zone are low. The Yellowjacket Zone has a relatively simple mineralogy that appears amenable to standard beneficiation technology.

Recent core examinations suggest a CRD model with Ag-Pb-Zn associated with pervasive silica alteration of the host carbonates. Generally, the Yellowjacket Zone lies 20 to 50 metres into the hanging wall (northeast) from – and sub parallel to – the Main Zone. However, Yellowjacket-style mineralization is also observed to occur in the immediate hanging wall of the Main Zone. The Yellowjacket Zone is composed of multiple sub zones separated by alternating units of limestone and phyllites. The mineralization and pervasive silica alteration are strongest in the core contact and gradually reduces into each lithology. The Yellowjacket Zone is composed of a disseminated to patchy massive red and honey-coloured (zinc-rich, lowiron) sphalerite with minor disseminated and hairline veinlets of galena and elevated silver values. The mineralization appears to be confined to favourable carbonate rocks that have been folded into a recumbent overturned anticline straddled by phyllite. Grade and thickness increase towards the interpreted fold hinge. Fluorite is commonly present while barite is absent in the Yellowjacket Zone.

9.0 EXPLORATION

No exploration has been conducted on the Revel Ridge Property since 2012 other than soil sampling and the drilling program carried out by Rokmaster in 2020, the results of which (as discussed in Section 10) were not available for inclusion in this Report.

Huakan conducted underground drifting and drilling between 2010 and 2012. Details of the 2010 to 2012 drill programs undertaken by Huakan are provided in Section 10.1 and all other exploration is discussed in Section 6.0.

10.0 DRILLING

In 2020, Rokmaster Resources completed 6,332 m of drilling in 20 NQ drillholes collared from the Revel Ridge underground workings. Analytical results from those drillholes are largely outstanding at the effective date of this Technical Report. Therefore, the 2020 drillholes, and the results of those drillholes, are excluded from this Technical Report.

The following section describes drilling carried out by Huakan and is summarized from the Technical Report by Puritch et al. (2020).

Table 10.1 provides a summary of all the drilling carried out at the Property, including the most recent work undertaken by Huakan. A total of 22,114 metres were drilled prior to Merit/Huakan taking control, who then drilled an additional 18,962 metres. All previous drilling and exploration work is summarized in Section 6.0.

Year	Drillholes	Total Metres	Company
1962-1967	UG DDHs	183.0	Westairs Mines Ltd.
1983-1984	65 UG DDHs	2,640.0	BP Selco Ltd.
1987-1988	20 UG DDHs	1,914.0	Pan American Minerals
1988-1989	32 UG DDHs	2,985.0	Equinox Resources Ltd.
1990-1991	50 UG DDHs	13,889.0	Equinox Resources Ltd.
	27 Surface DDHs		Cheni Gold Mines Ltd.
1997	3 UG DDHs	503.0	WeyminMining Corp.
2006	2-4? UG DDHs	undisclosed	BACTECH Mining Corp.
2007	9 Surface DDHs	1,363.0	Merit Mining Corp.
2010-2011	60UG DDHs	7,874.0	Merit/Huakan International
2011-2012	45 UG DDHs	9,724.9	Huakan International
	TOTAL	41,075.9

Table 10.1 Drill Programs Summary

10.1 HUAKAN DIAMOND DRILLING

The 2010 to 2012 drilling undertaken by Huakan has aided in the verification of the pre-2000 drilling performed at the Property and has been summarized below.

In November 2010, Merit/Huakan commenced a Phase 1, 2,000 metre underground drill program with the aim to verify historic drilling and to broaden the known mineralization of the Main Zone and provide support for a National Instrument 43-101 Mineral Resource Estimate. The Phase 1 program, totalling 7,874 metres over 60 BQTW (thin wall) core holes, was focused only on the Main Zone and was completed by early February 2011.

From June to August 2012, Huakan completed a 9,725 metre, 45-hole, underground drill program. The program's objective was to extend the 2011 Mineral Resource area and include infill drilling on the Yellowjacket Zone.

In the winter of 2011 and spring of 2012, Huakan completed 450 metres of track drifting, extending the length of the workings to the southwest and provide drill bays to drill the southeastern edge of the Main Zone.

In May 2012, Huakan commenced, and by mid-June 2012 had completed a 45-hole, 9,725 metre, underground diamond drill hole program to expand the Main Zone Indicated Mineral Resource and to infill the Yellowjacket Zone. The program was successful in intersecting similar grade and thickness of mineralization as previous nearby holes. The entire 2012 program tested six target areas, A through F, which are identified on a longitudinal section in Figure 10.1. Hole density or spacing in this campaign averaged 60 metre centres.

Eleven holes in Area A covered a 200-metre long by 130-metre down dip area of the Main Zone. Intercepts ranged between 0.56 metres and 8.48 metres of typical Main Zone mineralization. The length weighted average gold grade for all intercepts in this area was 5.55 g/t Au. Multiple zones were encountered in some drill holes with highlights tabulated in Table 10.2 and Table 10.3.

Drill HoleID	From (m)	To(m)	Length(m)	Au(g/t)	Ag(g/t)	Pb(%)	Zn(%)
DDH12-08	171.75	174.74	2.99	2.88	73.29	3.63	9.35
DDH12-09	192.38	193.61	1.23	4.13	71.88	3.03	2.68
DDH12-09	206.55	208.11	1.56	1.94	20.94	0.21	0.06
DDH12-10	207.70	216.18	8.48	9.41	101.39	2.17	4.31
DDH12-11	207.58	211.85	4.27	3.61	89.77	2.00	1.66
DDH12-26	212.35	214.27	1.92	7.36	191.31	6.58	3.97
DDH12-27	225.54	226.34	0.80	4.44	72.16	2.77	5.32
DDH12-27	227.75	228.31	0.56	2.18	24.30	1.19	8.06
DDH12-27	238.90	239.57	0.67	3.01	22.10	1.25	0.42
DDH12-29	221.52	222.49	0.97	4.64	93.32	2.91	8.81
DDH12-29	225.80	227.37	1.57	8.32	90.61	2.82	5.89
DDH12-32	232.08	233.96	1.88	1.47	26.28	1.06	3.47
DDH12-32	246.31	248.11	1.80	2.93	35.74	0.38	1.12
DDH12-32	250.39	251.52	1.13	1.48	92.88	1.73	3.89
DDH12-35	222.29	224.70	2.41	8.51	54.02	2.12	2.69
DDH12-39	240.67	242.60	1.93	7.03	70.92	3.32	6.32
DDH12-43	262.10	265.42	3.32	7.88	65.20	2.17	2.84
DDH12-43	273.40	274.00	0.60	4.98	113.00	2.56	2.88

Table 10.2 Drill Programs Summary Main Zone Drill Highlights in Area A

Figure 10.1 Longitudinal Section Showing All Drill Hole Pierce Points and 2012 Drilling Areas

Table 10.3 Main Zone Drill Highlights in Area B

Drill Hole	From	To	Length	Au	Ag	Pb	Zn
ID	(m)	(m)	(m)	(g/t)	(g/t)	(%)	(%)
DDH12-17	199.04	200.53	1.49	2.12	65.19	3.78	7.19
DDH12-18	184.83	187.39	2.56	4.74	42.98	1.53	2.18
DDH12-18	196.15	197.82	1.67	28.20	27.15	0.43	0.36
DDH12-19	226.27	226.77	0.50	0.96	122.00	5.04	3.69
DDH12-20	180.22	181.93	1.71	3.28	26.39	0.99	1.85
DDH12-21	194.48	195.65	1.17	3.26	18.16	0.61	1.29
DDH12-21	199.01	199.62	0.61	3.63	6.80	0.03	0.03
DDH12-22	235.74	237.96	2.22	1.13	22.20	0.81	2.41
DDH12-22	244.95	245.64	0.69	11.30	71.30	1.38	1.13
DDH12-23	204.20	206.89	2.69	7.66	82.58	3.44	6.30
DDH12-24	211.44	213.55	2.11	9.79	22.96	0.78	1.81
DDH12-24	217.83	218.70	0.87	6.25	41.68	0.87	0.39
DDH12-25	218.69	220.68	1.99	6.84	36.46	1.51	2.20
DDH12-28	245.06	247.65	2.59	2.45	16.79	0.63	1.20
DDH12-28	256.45	258.44	1.99	4.32	9.08	0.32	0.10
DDH12-30	253.33	255.20	1.87	3.47	51.00	0.64	3.66

Eleven holes in Area B covered a 250-metre long by 120 metre down dip area of the Main Zone. In this area, there were often one or two Main Zone intercepts per hole with intercept widths between 0.50 metres and 2.69 metres of typical Main Zone mineralization. The length weighted average gold grade for all intercepts in this area was 5.66 g/t Au. Highlights of Main Zone interceptions for this area are tabulated in Table 10.4.

Ten holes were completed in Area C covering an area 150 metres long by 250 metres down dip on the Main Zone, in the far southeast end of the deposit. The Main Zone continued throughout this area. Main Zone intercepts ranged from 0.42 metres to 5.82 metres of typical Main Zone mineralization. The length weighted average gold grade for all intercepts (excluding DDH12-14) was 4.67 g/t Au. Highlights of Main Zone interceptions for this area are tabulated in Table 10.4.

Drill Hole	From	To	Length	Au	Ag	Pb	Zn
ID	(m)	(m)	(m)	(g/t)	(g/t)	(%)	(%)
DDH12-12	107.27	108.72	1.45	7.32	21.88	0.51	0.48
DDH12-13	102.07	102.77	0.70	3.58	11.00	0.24	0.05
DDH12-14	161.67	162.31	0.64	1.24	1.60	0.02	0.01
DDH12-15	103.30	109.12	5.82	4.20	14.15	0.31	0.51
DDH12-16	136.10	137.40	1.30	3.86	10.52	0.41	0.67
DDH12-16	141.93	142.95	1.02	4.34	13.44	0.11	0.05
DDH12-36	176.12	176.91	0.79	15.60	51.60	1.65	2.44
DDH12-37	104.96	105.38	0.42	6.13	156.00	3.97	4.05
DDH12-40	98.57	100.19	1.62	6.50	30.48	0.71	1.48
DDH12-41	129.24	133.30	4.06	2.88	36.79	1.56	1.40

Table 10.4 Drill Programs Summary Main Zone Drill Highlights in Area C

Three holes were completed in Area D covering an area 80 metres long by 100 metres down dip on the Main Zone. Intercept widths ranged from 1.39 metres to 1.41 metres of typical Main Zone mineralization. The length weighted average gold grade for all intercepts was 4.69 g/t Au. Highlights of Main Zone interceptions for this area are tabulated in Table 10.5.

Table 10.5 Drill Programs Summary Main Zone Drill Highlights in Area D

Drill HoleID	From(m)	To(m)	Length(m)	Au(g/t)	Ag(g/t)	Pb(%)	Zn(%)
DDH12-31	121.57	122.96	1.39	8.52	78.02	2.89	3.01
DDH12-33	125.88	127.29	1.41	3.65	41.91	2.59	2.53
DDH12-34	180.15	181.55	1.40	1.92	16.63	0.61	1.00

Three holes were completed in Area E covering an area 100 metres long by 80 metres down dip on the Main Zone. This area fills in the gap between Area F and the area of the previous Mineral Resource Estimate. The length weighted average gold grade for all intercepts in this area was 5.78 g/t Au. Highlights of Main Zone interceptions for this area are tabulated in Table 10.6.

Drill Hole	From	To	Length	Au	Ag	Pb	Zn
ID	(m)	(m)	(m)	(g/t)	(g/t)	(%)	(%)
DDH12-42	237.13	240.33	3.20	6.00	26.24	1.12	8.00
DDH12-44	260.87	262.48	1.61	6.17	14.68	1.04	4.86
DDH12-45	222.63	223.03	0.40	2.00	51.80	0.25	0.68
DDH12-45	233.99	234.35	0.36	7.26	35.70	0.30	0.07
DDH12-45	238.66	239.15	0.49	0.34	23.80	1.68	10.65
DDH12-45	242.98	245.08	2.10	3.68	31.40	1.74	2.31
DDH12-45	254.25	255.17	0.92	8.29	31.49	1.56	0.46
DDH12-45	262.36	262.63	0.27	28.90	80.10	2.34	2.50
DDH12-45	271.17	271.94	0.77	3.34	26.90	0.56	0.22

Table 10.6 Drill Programs Summary Main Zone Drill Highlights in Area E

Seven holes were completed in Area F covering an area 180 metres long by 180 metres downdip on the Main Zone. Main Zone intercept widths ranged from 0.93 metres to 6.65 metres of typical Main Zone mineralization. The length weighted average gold grade for all intercepts in this area was 5.59 g/t Au. These same seven holes intercepted multiple Yellowjacket (silver-lead-zinc) zones further up in the holes with intercept widths ranging from 1.04 metres to 3.25 metres (Table 10.8). Highlights of Main Zone interceptions for this area are tabulated in Table 10.7.

Table 10.7 Drill Programs Summary Main Zone Drill Highlights in Area F

Drill Hole	From	To	Length	Au	Ag	Pb	Zn
ID	(m)	(m)	(m)	(g/t)	(g/t)	(%)	(%)
DDH12-01	149.58	151.22	1.64	2.94	28.3	1.46	1.94
DDH12-02	182.57	183.95	1.38	2.69	41.7	1.84	1.82
DDH12-03	75.56	76.89	1.30	6.17	23.6	1.20	3.08
DDH12-04	139.90	146.55	6.65	5.23	34.6	0.90	2.18
DDH12-05	187.49	189.62	2.13	11.73	85.6	2.87	2.92
DDH12-06	198.68	199.61	0.93	5.28	62.1	3.08	2.77
DDH12-07	246.25	247.95	1.70	3.90	23.4	1.23	3.27

Table 10.8 Drill Programs Summary Yellowjacket Zone Drill Highlights in Area F

Drill HoleID	From(m)	To(m)	Length(m)	Ag(g/t)	Pb(%)	Zn(%)
DDH12-01	103.81	105.50	1.69	92.0	3.40	13.50
DDH12-01	115.86	117.46	1.60	98.0	4.90	8.80
DDH12-02	116.41	117.5	1.09	52.2	3.07	16.91
DDH12-02	129.16	132.28	3.12	99.4	4.75	14.51
DDH12-03	93.68	95.19	1.51	51.7	1.88	13.79
DDH12-03	103.89	105.92	2.03	75.9	3.91	14.53
DDH12-04	120.83	121.92	1.09	59.9	3.28	4.66
DDH12-05	97.83	100.43	2.60	58.0	2.20	3.50
DDH12-05	131.94	134.28	2.34	31.3	1.20	10.44

Drill Hole	From	To	Length	Ag	Pb	Zn
ID	(m)	(m)	(m)	(g/t)	(%)	(%)
DDH12-06	103.91	105.93	2.02	106.7	2.90	10.57
DDH12-06	130.16	133.62	3.46	33	1.80	4.80
DDH12-06	143.35	146.4	3.05	43.3	2.75	9.40
DDH12-06	158.05	159.09	1.04	78.3	3.05	3.29
DDH12-07	179.78	181.05	1.27	119.0	7.90	7.70
DDH12-07	197.2	200.45	3.25	100.0	5.50	2.60

Results from the 45-hole, 2012 drilling program were used to prepare an updated Mineral Resource Estimate by P&E (announced September 18, 2012) on the Main Zone and Yellowjacket Zone.

10.1.1 Collar Surveying

At the completion of the 2007 surface and the 2010/2011 and 2012 underground drill programs, drill hole collar locations of all holes were marked and surveyed by B.C. professional land surveyors.

10.1.2 Downhole Surveying

During the 2007 surface diamond drill program downhole surveys were carried out using an Easy-Shot tool, taking measurements at the bottom and midway for the first three holes. Due to a defective tool, the final three drill holes were tested by acid tests at the bottom of each hole.

Downhole surveying in the 2010/2011 and 2012 underground drill programs utilized the FLEXIT SmartToolTM Drill Hole Survey system. Measurements were taken every 30 metres down the hole, usually including a near to collar test as well as a near to bottom hole test. All azimuth readings taken during the downhole surveys had a magnetic declination factor of 17º added to them to give true azimuth readings for this region of British Columbia. Other data collected were dip angles recorded at the various downhole reading sites as well as magnetic susceptibility. Any strongly erroneous magnetic results put a small number of the azimuth readings in question, as to their reliability so were subsequently eliminated.

10.1.3 Core Recovery and Storage

Core recoveries throughout the 2010/2011 and 2012 underground J&L (now Revel Ridge) drill programs were normally >90% and often >95%. All drill core from both the 2007 and the 2010/2011 drill programs are securely stored on the Property, near the camp facility. All nonmineralized drill core from the 2012 drill program is securely stored on the Property, near the camp facility. All drill core that had mineralized intercepts from the 2012 drill program is securely stored in the warehouse just north of Revelstoke.

10.1.4 Drill Core Size and Orientation

All Huakan drilling discussed in this Technical Report section was done with wireline BQTW diamond core. True widths are approximately 75% of downhole intercept lengths. The mineralization dips NE at 50°.

10.1.5 Contractor

The 2007 diamond drill program was carried out by Elite Drilling Ltd. of Revelstoke, B.C. over the period October 23 to November 13, 2007.

DMAC Drilling of Aldergrove, BC was the drilling contractor for the 2010/2011 and 2012 program. For the 2010/2011 campaign, drilling took place between November 11, 2010 to January 31, 2011. For the 2012 campaign, drilling took place between May 6, 2012 and June 16, 2012. Drilling was carried out on two ten-hour shifts using two Hydracore drill rigs mounted on steel wheels, thus providing drill access to the tracked 830 main drift and crosscuts.

10.1.6 Comments

This Technical Report author is of the opinion that procedures undertaken by Huakan during the 2010 to 2012 drilling programs conform to standard industry practice and that there are no drilling, sampling or recovery factors that materially impact the reliability of the drill core data.

11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

Samples collected by Rokmaster in 2020 from its drill program, as well as from surface and underground, are excluded from this Technical Report, since many of the analytical results are outstanding as of the effective date of this Technical Report. No other sampling has been undertaken by Rokmaster. The following section is based on the Technical Report on the Property by Puritch et al. (2020).

11.1 SAMPLING METHOD AND APPROACH BY HUAKAN

A total of 956 split core samples from the Huakan 2010/2011 diamond drill program were collected by and analyzed for Huakan and a total of 895 split core samples from the 2012 diamond drill program. Sampling was carried out where visual sulphide concentrations were observed beyond non-mineralized host-rocks.

Sample intervals were generally less than 0.5 metres where stronger sulphide concentrations were observed and ranged between 0.5 to 1.0 m intervals. Occasional, narrower sample intervals ranged between 0.25 to 0.5 metres where intervals with massive veins were observed.

A total of 427 bulk density measurements (by wet immersion technique) were taken on site on the 2010/2011 drill core by competent company geological staff and a total of 86 on the 2012 core.

The following summary details the sampling procedures and steps taken during the 2010/2011 and 2012 drill programs by Huakan:

Drill core was first cleaned, organized and photographed;
Geotechnical logging was undertaken by a trained technician;
Drill core boxes were labelled using scribed aluminum tags;
Drill core logging and sample selection was performed by the site geologists;
In areas of Main Zone mineralization, sampling intervals were determined by similar sulphide abundance;
Sampling was carried out beyond the limits of the Main Zone sulphides both into barren hanging wall and footwall rocks;
Every 18th, 19th and 20th sample was designated as a duplicate, standard and blank, respectively. The duplicate sample was a fifty percent split of the sample preceding it;
Drill core was logged, sampled and stored on site. The logging geologist would place a colour crayon line along the desired sample cut to provide an even bisection of the core;
The drill core was cut in half, bisecting fabric or vein material evenly; and

• Technicians were instructed to place the same side of drill core back into the box for every sample and the other side into a labelled clean plastic sample bag that was then sealed using a zip tie.

11.2 CHAIN OF CUSTODY

Sample bags were placed in address-labelled rice bags, sealed with plastic zip ties and shipped from Revelstoke, B.C., by Greyhound Bus to Eco Tech Laboratory Ltd., ("Eco Tech") of Kamloops, BC. (later acquired and managed by ALS Minerals).

Sample shipment records were maintained. Records were also kept of sample preparation, analysis requested, and the person intended to receive the results.

Drill core sampling was carried out by use of a diamond blade core saw. The drill core sampler was highly experienced and sampling work was closely monitored by on-site core logging geologists.

No core samples were taken by an employee, officer, director or associate of Huakan.

11.3 SAMPLE PREPARATION AND ANALYSES

Analytical work for the 2010/2011 and 2012 drill programs was carried out by Eco Tech. Huakan has archived all of the original assay certificates for the 2007, 2010/2011 and 2012 drill programs.

Eco Tech's sample preparation and analysis procedures were as follows:

At the time of analysis, Eco Tech was registered for ISO 9001:2008 by KIWA International (TGA-ZM-13-96-00) for the provision of assay, geochemical and environmental analytical services. Eco Tech also participated in the annual Canadian Certified Reference Materials Project (CCRMP) and Geostats Pty., bi-annual round robin testing programs. Eco Tech operated an extensive quality assurance/quality control program, which covers all stages of the analytical process from sample preparation through to sample digestion and instrumental finish and reporting.
Samples (minimum sample size 250 g) are catalogued and logged into the sampletracking database, once received by the lab, and checked for spillage, general sample integrity and that samples matched the sample shipment requisition. The samples are transferred into a drying oven and dried. Rock samples are crushed by a Terminator jaw crusher to -10 mesh ensuring that 70% passes through a Tyler 10 mesh screen. This is verified each batch.
Re-split are taken every 35 samples using a riffle splitter and tested to ensure the homogeneity of the crushed material. A 250-gram sub sample of the crushed material is pulverized on a ring mill pulverizer, each batch ensuring that 85% passes through a 200 mesh screen. The sub sample is rolled, homogenized and bagged in a pre-numbered

bag. A barren gravel blank is prepared before each job in the sample prep to be analyzed for trace contamination along with the actual samples.

Samples analyzed for gold (30-gram sample size) are fire assayed along with certified reference materials ("CRMs" or "standards") using appropriate fluxes. The flux used is pre-mixed, purchased from Anachemia and contains Cookson Granular Litharge (Silver and Gold Free). The ratios are 66% Litharge, 24% Sodium Carbonate, 2.7% Borax, 7.3% Silica. (These charges may be adjusted with borax or silica based on the sample). Flux weight per fusion is 120 g. Purified Silver Nitrate is used for inquartation. The resultant doré bead is parted and digested with aqua regia and then analyzed on an atomic absorption instrument (Perkin Elmer/Thermo S-Series AA instrument). Gold detection limit on AA is 0.03-100 g/t. Any gold samples over 100 g/t are run using a gravimetric analysis protocol. Each batch submitted is fire assayed as a batch.
Appropriate standards and repeat/re-split samples (Quality Control Components) accompany the samples on the data sheet for quality control assessment. For 30 element ICP, a 0.5-gram sample is digested with a 3:1:2 (HCl:HNO3:H2O) for 90 minutes in a water bath at 95°C. The sample is then diluted to 10 ml with water. All solutions used during the digestion process contain beryllium, which acts as an internal standard for the ICP run. The sample is analyzed on a Thermo Scientific IRIS Intrepid II XSP/ICAP 6000 Series ICP unit. CRMs are used to check the performance of the machine and to ensure that proper digestion occurred in the wet lab. QC samples are run along with the client samples to ensure no machine drift or instrumentation issues occurred during the run procedure. Repeat samples (every batch of 10 or less) and re-splits (every batch of 35 or less) are also run to ensure proper weighing and digestion occurred. Results are printed along with accompanying quality control data (repeats, re-splits and standards). Any of the base metal elements that are over limit, Ag >30 g/t, Cu, Pb, Zn >1.0%) are run as an assay.

It is the Author's opinion that the sampling preparation, security and analytical procedures employed by Huakan were satisfactory to support a Mineral Resource Estimate.

11.4 HUAKAN QUALITY ASSURANCE/QUALITY CONTROL

11.4.1 Certified Reference Material

Huakan geologists routinely inserted Certified Reference Material (CRM) and blanks into the sample stream during the 2010/2011 and 2012 drill programs. The CRMs and blanks were obtained from CDN Resource Laboratories of Langley, B.C.

The CDN standards were CDN-ME-7 and CDN-ME-11 in the 2010/2011 program and CDN-ME-8 and CDN-ME-11 in the 2012 program. CRMs were inserted into the sample stream at a rate of 1 in 20 by the project geologists. CRMs are inserted regularly into batches of samples sent to the lab for analysis in order to monitor the accuracy (lack of bias) of the lab results.

A total of 36 data points was available for the CDN-ME-7 CRM and 29 for the CDN-ME-11 CRM, for the 2010/2011 program. Both CRMs were certified for gold, silver, lead and zinc and both performed very well, with all data points falling within +/- two standard deviations from the mean certified value.

A total of 15 data points was available for the CDN-ME-8 CRM and 22 for the CDN-ME-11 CRM, for the 2012 program. Both the CDN-ME-8 and CDN-ME-11 CRMs were certified for gold, silver, lead and zinc. Both CRMs performed well, with the majority of data points falling within +/- two standard deviations from the mean certified value.

The majority of data points for the CDN-ME-8 CRM fell within +/- two standard deviations from the mean certified value. All data points for zinc fell within + two standard deviations from the mean, displaying a slight high bias. All data for lead fell within – three standard deviations from the mean, displaying a slight low bias. For both gold and silver, one data point fell above + three standard deviations from the mean, and the remaining data points fell within three standard deviations from the mean certified value. A slight high bias was also noted for gold and silver for this CRM.

The majority of data points for the CDN-ME-11 CRM fell within +/- two standard deviations from the mean certified value. All data points for zinc fell within + two standard deviations from the mean, displaying a slight high bias. All data points for lead fell within – three standard deviations from the mean, displaying a slight low bias. For both gold and silver, one data point fell above + three standard deviations from the mean, and the remaining data points fell within three standard deviations from the mean certified value. A slight high bias was also noted for gold and silver for this CRM.

11.4.1.1 Blanks

Huakan purchased blanks consisting of pulverized river rock (predominantly granite) from CDN Resource Laboratories Ltd., of Langley, BC, for use in the 2010 to 2012 drilling programs. CDN Resource Laboratories Ltd.'s, assaying of the blank material found it to contain <0.01 g/t Au. It is not a common practice to use pulverized blank material since it will not pass through the crushers and splitters where most thereby limiting contamination detection. Blanks were inserted into the sample stream at a rate of 1 in 20.

All data points for gold, for 2010/2011 and 2012, and silver, for 2010/2011, were below the upper threshold of three times the detection for the element in question, which was the upper threshold set for monitoring blank results. There were four outliers (out of a total of 40 data points) observed for silver for the 2012 data. For the 2010/2011 drilling program, lead returned an average value of 0.002% with a standard deviation of 0.0006. Zinc returned an average value of 0.005% with a standard deviation of 0.0005. For the 2012 drilling program, lead returned an average value of 0.001% with a standard deviation of 0.0008. Zinc returned an average value of 0.006% with a standard deviation of 0.0014. All results indicate no contamination present at the analytical level.

11.4.1.2 Duplicate Sampling Program

Field duplicates were implemented as part of the Quality Assurance/Quality Control (QA/QC) sampling protocol for both the 2010/11 and 2012 drilling programs, in order to quantify precision (reproducibility) of analytical results at the field level.

Drill core duplicates were inserted into the sample stream at a rate of 1 in 20. A duplicate sample consisted of a 50% split of the numbered sample interval immediately preceding the duplicate sample.

In addition, P&E examined the laboratory coarse reject duplicates and pulp duplicates for gold, silver, lead and zinc for the 2010/2011 program. The coarse reject data set contained on average 27 pairs, and the pulp data set contained 239 pairs for gold, 95 pairs for silver, 100 pairs for lead and 104 pairs for zinc.

Simple scatter graphs for all elements were plotted for all available data. Precision was observed to improve steadily from the core duplicates through to the pulp duplicates. The precision at the pulp duplicate level for all four metals was excellent, with a 1:1 ratio.

The author considers the data to be of good quality and satisfactory for use in a Mineral Resource Estimate.

12.0 DATA VERIFICATION

The following section is based on the Technical Report on the Property by Puritch et al. (2020).

12.1 2010/2011 SITE VISIT AND INDEPENDENT SAMPLING

The Revel Ridge Property was visited by Mr. Fred Brown, P.Geo., of P&E on December 17, 2010. Data verification sampling was done on diamond drill core, with 18 samples distributed in 18 holes collected for assay. These samples were collected from both the current drill program as well as from a number of the historic (1991 and earlier) drill holes. An attempt was made to sample intervals from a variety of low and high-grade material. The chosen sample intervals were then sampled by taking complete sections of the remaining half-split core. The samples were then documented, bagged, and sealed with packing tape and were delivered by Mr. Brown to ALS Minerals (formerly referred to as ALS Chemex), 2103 Dollarton Highway in North Vancouver for analysis.

ALS Minerals has developed and implemented strategically designed processes and a global quality management system at each of its locations that meets all requirements of International Standards ISO/IEC 17025:2017 and ISO 9001:2015. All ALS geochemical hub laboratories are accredited to ISO/IEC 17025:2017 for specific analytical procedures.

The ALS quality program includes quality control steps through sample preparation and analysis, inter-laboratory test programs, and regular internal audits. It is an integral part of dayto-day activities, involves all levels of ALS staff and is monitored at top management levels.

At no time, prior to the time of sampling, were any employees or other associates of Huakan advised as to the location or identification of any of the samples to be collected. A comparison of the P&E independent sample verification results versus the original assay results for gold, silver, lead and zinc can be seen in Figure 12.1 to Figure 12.4.

Figure 12.1 2010/2011 P&E Verification Samples for Gold

Figure 12.2 2010/2011 P&E Verification Samples for Silver

Figure 12.3 2010/2011 P&E Verification Samples for Lead

Figure 12.4 2010/2011 P&E Verification Samples for Zinc

12.2 2012 SITE VISIT AND INDEPENDENT SAMPLING

The Property was visited by Mr. Richard Routledge, P.Geo., of P&E from June 13 to 14, 2012. Data verification sampling was done on diamond drill core, with 26 samples from 10 drill holes collected for assay. An attempt was made to sample intervals from a variety of low and highgrade material. The chosen sample intervals were then sampled by taking complete sections of the remaining half-split core. The samples were then documented, bagged, and sealed with packing tape and were delivered by Mr. Routledge to AGAT Laboratories (AGAT) in Mississauga, ON, for analysis.

Samples at AGAT were analyzed for gold by fire assay with ICP-OES or gravimetric finish; silver, lead and zinc by aqua regia digest with ICP-OES finish; and lead and zinc samples exceeding 10,000 ppm were further analyzed using a Sodium Peroxide Fusion method with ICP-OES finish. Specific gravities were also determined on all 26 of the samples.

AGAT has developed and implemented at each of its locations a Quality Management System (QMS) designed to ensure the production of consistently reliable data. The system covers all laboratory activities and takes into consideration the requirements of ISO standards.

AGAT maintains ISO registrations and accreditations. ISO registration and accreditation provide independent verification that a QMS is in operation at the location in question. Most AGAT laboratories are registered or are pending registration to ISO 9001:2000.

Figure 12.5 2012 P&E Verification Samples for Gold

Figure 12.6 2012 P&E Verification Samples for Silver

Figure 12.7 2012 P&E Verification Samples for Lead

Figure 12.8 2012 P&E Verification Samples for Zinc

Based upon the evaluation of the QA/QC program undertaken by Huakan, as well as P&E's due diligence sampling, it is the Author's opinion that the results are suitable for use in the current Mineral Resource Estimate.

12.3 DATABASE VERIFICATION

The Author undertook a verification review of the supplied database whereby independently acquired laboratory assay certificates were compared to constrained assays within the Mineral Resource wireframes. 100% of all constrained assays from 2010 onwards were checked and seven errors were found and corrected. Prior to 2010 selected assays were checked and 25 corrections were made. There were no limitations on the Author's ability to conduct satisfactory data verification.

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 INTRODUCTION

The conceptual metallurgical flowsheet and design criteria used for the PEA is based mainly on a laboratory testwork program that was conducted under the direction of F. Wright Consulting Inc. on behalf of Huakan International Mining Inc (Huakan). Most of the metallurgical testwork was undertaken between 2011 and 2013 by Inspectorate Exploration and Mining Services Ltd., (Inspectorate) of Richmond, BC, with specific procedures completed by other laboratories including Hazen Research in Golden CO., and SGS in Lakefield ON.

The 2011-13 metallurgical testwork program used mineralized composite samples selected from both the Main Zone and the Yellowjacket Zone. The principal valuable metals contained in these two zones include gold (Au), silver (Ag), zinc (Zn) and lead (Pb), roughly in order of economic importance. The Main Zone constitutes most of the tonnage within the Revel Ridge Project and has a higher in-situ value due to the higher gold content. The Yellowjacket Zone has lower tonnage and only minor gold credits, but higher Ag, Zn and Pb content and typically tends to be easier to process than the Main Zone.

Some additional development testwork using samples of Main Zone mineralization has recently been undertaken by Base Metallurgical Laboratories Ltd. (Base Met) of Kamloops, BC, under the direction of Canenco Consulting Corp. (Canenco). The work completed and reported during 2020 includes gravity separation and flotation tests.

13.1.1 Historical Testwork

Numerous mineralogical and metallurgical studies have been carried out using mineralized Main Zone samples. The metallurgy studies and assessments date from 1982 through to 2005 with particular effort made throughout the 1980's and 1990's combined with exploration programs carried out by BP Canada (Selco) Ltd., Pan American Minerals Corporation, Equinox Resources Ltd., Cheni Gold Mines Ltd. (CGM) and Weymin Mining Corporation. A scoping study was produced in 1998 by H.A. Simons which used the results from metallurgical studies undertaken on a bulk sample collected by Weymin. Subsequently, PRA (now Bureau Veritas Minerals) of Richmond BC, completed a metallurgical study in 2005 on behalf of BacTech.

The majority of the historic testwork was focused on producing saleable lead and zinc concentrates by various flotation separation techniques. In addition, some tests were carried out on the refractory gold arsenopyrite concentrates where Cashman (chloride leach), Redox (nitrate leach), batch roasting, pressure oxidation and bio-oxidation techniques have been investigated, followed by cyanidation, to recover the precious metals.

Yellowjacket Zone mineralization was not used in the historic testwork studies.

One series of flotation testwork of note was completed by Bacon, Donaldson and Associates Ltd. In 1991 using a composite sample of heavy media sinks material provided by SGS Lakefield. Mineralogical data suggests that a high proportion of gold is associated with arsenopyrite and therefore this program of testwork looked at selectively floating pyrite and producing a gold rich arsenopyrite product as well as the lead and zinc concentrates. During non-optimized batch tests, an arsenopyrite concentrate containing around 30% As and 35 g/t Au was produced by the laboratory.

These results were duplicated in 2020 by Base Met under the direction of Canenco.

13.1.2 Recent Mineral Processing Testwork (2011-2013)

Metallurgical testing was completed on mineralized samples representing the Yellowjacket and Main Zones. This testwork program was designed to build on historical work conducted for the project primarily during the 1980's and 1990's. The program comprised various mineral processing and hydro-metallurgical procedures including comminution testing, heavy media separation (HMS) and differential flotation to produce separate Pb and Zn concentrates, and a gold containing pyrite/arsenopyrite concentrate. Gold bulk flotation concentrate was separately subjected to bioleaching and pressure oxidation (POX) procedures prior to cyanidation.

13.1.2.1 Metallurgical Samples

Samples from the Main Zone were selected, abstracted and shipped by Huakan to Inspectorate in June, 2011. These samples comprised seven bulk underground samples (six mineralized and one dilution rock), each consisting of minus 150 mm (-6") rock and weighing a total of 7.3 t. The six mineralized samples with varying grades were spaced along a 440-metre strike length of the deposit.

The samples representing the Yellowjacket Zone arrived at Inspectorate in October, 2012. The material included 77 kg of split drill core and 43 kg of minus 3.36 mm (-6 Tyler mesh) assay rejects.

The results of chemical analyses undertaken on the samples of the seven Main Zone composite samples are presented in Table 13.1. The analytical results for the master composite (JL1), which is a blend of the six mineralized composites selected to represent the Main Zone resource, particularly the average expected lead and zinc grade. The most recent 2020 average mineral resource grades are also included for comparison purposes.

The detailed assay data also showed the mercury content of the mineralized samples ranged from 4.5 to 21 ppm and that virtually all of the sulphur was present as sulphide sulphur.

Composite#	Au	Ag	Zn	Pb	As	Sb	STOT
	(g/t)	(g/t)	(%)	(%)	(%)	(ppm)	(%)
1	10.3	69.0	2.19	2.48	9.41	1663	12.0
2	12.6	80.0	2.39	2.30	8.83	3729	16.2
3	15.1	119	8.84	4.92	8.78	1419	16.9
4	13.3	120	7.42	3.61	8.44	2237	31.9
5	11.1	24.7	2.61	0.72	8.74	332	14.5
6	6.98	72.5	6.14	4.03	6.20	2295	12.9
7	0.28	11.6	2.14	0.58	0.50	50	0.5
JL1	6.99	57.6	3.90	2.45	5.90	1616	11.4
M&I Resources	5.59	53.4	3.43	1.87	-	-	-

Table 13.1 Main Zone Metallurgical Composite Sample Chemical Analyses

The Yellowjacket Zone samples originated from split drill core, as well as from minus 6 mesh assay rejects from earlier exploration programs. The drill core sample composites were labeled as Composite 1 (DC07), Composite 2 (DC12), Composite 3 (DC-LG) and Composite 4 was made up from the assay rejects. Composite 5 was a blend of DC07 and DC12. A summary of the sample analyses is provided in Table 13.2.

Table 13.2 Yellow Jacket Metallurgical Sample Chemical Analyses

Composite#	Au(g/t)	Ag(g/t)	Zn(%)	Pb(%)	As(%)	Hg(ppm)	STOT(%)
1 (DC07)	0.14	58.6	7.69	2.32	156	104	5.14
2 (DC12)	0.35	71.6	8.46	3.21	1435	115	6.28
3 (DC-LG)	0.24	52.6	4.67	1.68	179	65	3.12
4 (rejects)	0.15	73.1	9.04	3.21	503	117	6.44
5 (DC7&12)	0.13	62.1	7.96	2.68	954	102	5.57
M&I Resources	0.09	62.6	9.93	2.60	-	-	-

13.1.2.2 Mineralogy

The historic mineralogical studies and investigations showed that the Main Zone is a fine grained massive polymetallic mineralized body with complex mineralogy. These studies suggested a high degree of variability in the various samples although the major sulphide minerals present included pyrite, arsenopyrite, galena, freibergite, sphalerite, chalcopyrite and minor sulphosalt minerals such as bournonite (PbCuSbS3) and boulangerite (Pb5Sb4S11).

Mineral particle sizes and associations were highly variable although typically iron sulphides were discreet from the galena and sphalerite that were more finely intergrown. The finegrained mineralization will require very fine primary grinding, 80% passing (P80) of around 40 μm followed by rougher/scavenger concentrate regrinding to achieve adequate liberation.

Gold in the Main Zone typically appears in two modes. The majority (80%-90%) occurs as solid solution within the arsenopyrite matrix. The second mode of gold occurrence (10%-20%)

is as fracture filling/veinlets in arsenopyrite or as coarse grains locked in other sulphides of gangue. This secondary gold occurrence liberates on grinding, floats with the lead concentrate and is cyanide leachable. The majority of the silver is present as freibergite and is in solid solution with the lead minerals. Impurities are present either due to mineralogical makeup, or as less than 10 μm locked particles. Total metal recoveries to sulphide concentrates are expected to be high; however, good recoveries of lead and zinc into their respective separate concentrates will be limited because of cross-contamination due to fine particle locking.

No mineralogical studies have been completed using representative Yellowjacket Zone samples, however, based on the mineral samples used for the metallurgical program, the Yellowjacket mineralization tends to be significantly less complex than the Main Zone, with higher average Ag, Pb, Zn head grades and less intimate association of the galena and sphalerite. Typically, there is only minor pyrite, arsenopyrite, and sulphosalt minerals present resulting in reduced deleterious elements compared to the Main Zone.

13.1.2.3 Comminution Tests

A preliminary program of comminution tests comprised abrasion tests, crushing tests and standard grinding work indices determinations of the various samples. Bond crushing work index and Bond rod mill work index work was completed at Hazen Research in Colorado. Abrasion testing was performed by a contractor under Hazen's direction. Bond ball mill work index tests and the Starkey procedure were used by Inspectorate to provide preliminary ball mill and SAG mill design data.

The crushing work index (CWi) test results for composite 2, 3, and 4 ranged from 9.7 to 12.7 kWh/t.

The abrasion test result on the Main Zone master composite JL1 using the Pennsylvania crusher method was 0.2402 g. A rod mill work index of 12.9 kWh/t was reported for the same sample.

The SAG tests estimated a SAG Pinion energy requirement of 6.38 kWh/t for composite JL1.

The results from the Inspectorate Bond ball mill work index tests using a closing sieve size of 105 microns are summarized in Table 13.3. These tests used feed samples of Main Zone composites 1 to 6 but heavy media test "sinks" for Main Zone composite JL1 and Yellowjacket composite 5.

The results from the preliminary comminution tests suggest that both the Main Zone and Yellowjacket Zone mineralization is relatively soft. Although there was a high degree of variability in the Main Zone crusher index results, the Bond ball mill index test results were relatively consistent.

Composite ID	Index(kWh/t)
MZ-1	9.6
MZ-2	9.7
MZ-3	9.1
MZ-4	8.8
MZ-5	9.7
MZ-6	9.7
MZ-JL1 sinks	9.5
YJZ-5sinks	9.8

Table 13.3 Bond Ball Mill Index Test Results

YJZ=Yellowjacket Zone

13.1.2.4 Heavy Media Separation

Sink-float tests were performed by Inspectorate using the heavy liquid tetrabromoethane (TBE), with the concentration adjusted to achieve the desired specific gravity (SG) of the liquid media. The initial test work was performed on the Main Zone master composite (JL1) with the media SG at 2.72, 2.82 and 2.95 using two crush sizes, 25 mm (1") and 50 mm (2"). The heavy liquid test feed material was screened at 10 mesh (2 mm) to remove fines.

Following the initial scoping tests a bulk test (HLS 4) was performed at a media SG 2.85 using 10 kg batches of Composite JL1 to produce material for flotation feed. The solids SG of the combined float material was 2.77 and for the combined sink product 3.90. The minus 10 mesh material removed from the 200 kg bulk feed sample was 6.2% by weight.

Composites 1 to 6 were tested using a crush size of 50 mm and SG of 2.85 as part of the variability testwork program. A single test was undertaken using Yellowjacket mineralization (Composite 5, DC07&12) using a crush size of 18.7 mm (¾") and media SG of 2.85. The results for the heavy liquid tests undertaken using Main Zone composites and the single Yellowjacket test are presented in Table 13.4.

	Sample	Feed Size	Media			Sink Distribution (%)
Test ID	ID	(mm)	SG	Weight	Au	Ag	Pb	Zn
HLS1	JL1	50	2.95	52.3	97.5	93.6	92.9	86.7
HLS1	JL1	50	2.82	63.9	98.6	96.9	97.3	94.6
HLS1	JL1	50	2.72	74.4	99.6	98.4	98.8	98.0
HLS2	JL1	25	2.95	48.7	97.4	93.9	94.1	89.8
HLS2	JL1	25	2.82	55.6	98.6	96.4	96.1	92.5
HLS2	JL1	25	2.72	74.5	99.9	99.2	98.8	97.6
HLS3	JL1	50	2.96	55.5	98.7	95.4	94.0	86.4
HLS4*	JL1	50	2.85	59.0	99.3	97.2	96.6	91.5
HLS5	MZ-1	50	2.85	78.0	99.6	95.0	99.5	99.3

Table 13.4 Heavy Liquid Test Results – Main Zone

MZ= Main Zone

	Sample	Feed Size	Media	Sink Distribution (%)
Test ID	ID	(mm)	SG	Weight	Au	Ag	Pb	Zn
HLS6	MZ-2	50	2.85	86.2	99.3	99.0	99.1	97.3
HLS7	MZ-3	50	2.85	80.0	99.9	99.5	99.5	99.8
HLS8	MZ-4	50	2.85	92.8	99.6	99.9	100.0	99.8
HLS9	MZ-5	50	2.85	58.4	99.6	98.3	99.1	99.5
HLS10	MZ-6	50	2.85	77.0	99.1	98.8	99.2	96.1
YJ-HLS1	YJ-5	18.7	2.85	42.1	80.7	83.2	83.1	87.4

* Bulk test to produce flotation feed.

The results show the Main Zone sub-composites used in the master composites had excellent metal recovery, but that the weight distribution between sink and float portions is variable. This is dependent on the proportion of sulphide minerals in the sample. For the Main Zone samples the sulphides would appear to respond as massive to semi-massive with minor losses to the HMS float. The relationship of sample sulphur grade to sinks weight recovery is presented in Figure 13.1. These data include tests HLS4 to HLS10 for the different Main Zone composites from which a curve is fitted so that an approximate sink to float distribution can be estimated from the proportion of sulphides in the feed.

Figure 13.1 Main Zone Heavy Liquid Test Results for Different Sample Composites

The performance of the single Yellowjacket test was significantly worse than the Main Zone tests, possibly due to the more disseminated nature of the sulphide mineralization.

13.1.2.5 Flotation – Main Zone

A series of open circuit optimization flotation tests were undertaken using the sinks from the heavy liquid bulk sample test (Main Zone master composite JL1) combined with the -10 mesh screen fines. Historic test work had shown a fine primary grind is required for the separation

of lead and zinc, so initial tests were performed using a grind P80 range of 35 to 65 microns. The results from these grind tests showed only minor improvement of Zn and Pb recoveries, but a significant trend of higher Au and Ag recoveries, with finer grinding. The remaining Main Zone flotation testwork work targeted a primary grind P80 in the 30 to 40-micron range.

Based on the results from historical flotation testwork, the 2013 flotation optimization program considered the following:

1. Differential lead, zinc and iron/arsenic sulphide flotation.
1. Combined bulk flotation of galena and sphalerite followed by separation of lead and zinc during cleaning.
1. Lead float followed by combined bulk flotation of sphalerite with iron/arsenic sulphide minerals followed by subsequent depression of the iron sulphide during cleaning of sphalerite.

Separation of arsenopyrite and pyrite to produce an upgraded gold concentrate was identified during the 2013 test program as an option worth considering in the future, as the majority of the gold was considered to be associated with arsenopyrite. This flowsheet option was revisited in the 2020 flotation test program managed by Canenco.

Following a number of open circuit roughing and cleaning tests that investigated the three options, two locked cycle tests (LCT) were completed using Composite JL1. The first LCT was based on the differential flotation of lead, zinc and iron/arsenic sulphides (see Figure 13.2) and the second LCT comprised the bulk flotation of sphalerite with iron/arsenic sulphide minerals.

Figure 13.2 Main Zone Locked Cycle Flotation Test FLC1 Flowsheet

Source: F. Wright, Appendix MZ-6, Locked Cycle Flotation Tests FLC 1, 2, 3, 4, 4A, 5.

The results of the two Main Zone locked cycle flotation tests are summarized in Table 13.5 and Table 13.6.

				Grades				Metal Distribution (%)
Product	Wt.(%)	Au(g/t)	Ag(g/t)	Zn(%)	Pb(%)	As(%)	S(%)	Au	Ag	Zn	Pb
Pb con	2.9	29.8	720	8.7	58.4	3.1	18.5	8.1	33.7	5.2	57
Zn con	4.1	1.4	40.9	61.1	1.2	0.49	32.5	0.5	2.7	51.4	1.7
Au con	62.4	15.5	62.9	3.3	1.9	13.5	24.1	90.0	62.6	42.4	40
Tail	30.5	0.49	2.0	0.16	0.15	0.99	1.4	1.4	1.0	1.0	1.5
Feed (calc.)	100.0	10.8	62.6	4.90	2.99	8.83	17.4	100.0	100.0	100.0	100.0

Table 13.5 Summary of Flotation Test Results – Main Zone Composite JL1 LCT1

Table 13.6 Summary of Flotation Test Results – Main Zone Composite JL1 LCT2

				Grades				Metal Distribution (%)
Product	Wt.(%)	Au(g/t)	Ag(g/t)	Zn(%)	Pb(%)	As(%)	S(%)	Au	Ag	Zn	Pb
Pb con	2.8	72.6	875	11.8	44.5	3.2	20.2	16.5	49.6	5.8	52.4
Zn con	9.2	4.9	114	54.8	5.0	1.4	31.2	3.6	20.9	4.4	19.2
Au con	48.6	20.2	28.4	0.6	1.3	18.3	29.2	78.9	27.7	5.0	25.7
Tail	39.5	0.32	2.18	0.13	0.17	0.20	0.85	1.0	1.7	0.9	2.7
Feed (calc.)	100.0	12.4	49.8	5.69	2.40	9.20	17.9	100.0	100.0	100.0	100.0

Following a review of the two LCT test results it was decided to undertake additional variability locked cycle flotation tests using Main Zone Composites 1, 3 and 6, which represented the range of expected typical lead and zinc head grades for the Main Zone. These tests were undertaken using feed composite samples not heavy liquid sink material. The threeproduct differential flowsheet (Option 1) was selected for this variability testwork. A summary of the Main Zone variability LCT results is presented in Table 13.7 (over).

The calculated head grades for all the LCTs undertaken on the main zone composites are provided in Table 13.8 (over). Generally, separation and recovery of lead and zinc appears reasonable although elevated zinc can remain in the lead concentrate for some composites, which may reduce payables from some smelters. With only one exception, the lead grades remained above 50%, and zinc above 54%, in their respective concentrates.

These variability locked cycle flotation tests indicated that the two base metal concentrates can contain a fairly wide range of other elements, both valuable and deleterious. Typically, smelter terms precious metals credits are higher in the lead concentrate compared to the zinc concentrate and favorably, these tests suggest that most of the precious metals recovered to the base metal products report to the lead concentrate. Gold deportment to the lead concentrate varied from 6% to 22%, averaging about 11% but with less than 2% reporting with zinc. Silver recovery to the lead concentrate was between 34% and 80%, averaging around 60%, with about 10% of the silver directed to the zinc concentrate.

						Lead Concentrate
Test	Comp.		Lead		Gold		Silver			Grades
ID	ID	Wt.	Grade	Rec.	Grade	Rec.	Grade	Rec.	Zn	As	Sb
		(%)	(%)	(%)	(g/t)	(%)	(g/t)	(%)	(%)	(%)	(%)
FLC1	JL1	2.9	58.9	57.2	29.8	8.1	720	33.7	8.7	3.0	-
FLC3	1	2.3	58.8	67.8	22.7	5.8	723	37.1	3.4	4.2	4.3
FLC4A	3	7.9	57.4	90.1	13.6	6.6	721	79.2	8.5	3.0	0.9
FLC5	6	7.8	46.6	90.0	17.9	22.5	656	80.1	14.4	3.9	2.5
Average (all)		5.2	55.4	76.3	1.0	10.8	705	57.5	8.8	3.5	2.6
Avg. FLC3, 4A, 5		6.0	54.3	82.6	18.1	11.6	700	65.5	8.8	3.7	2.6
						Zinc Concentrate
Test	Comp.		Zinc		Gold		Silver			Grades
ID	ID	Wt.(%)	Grade	Rec.	Grade	Rec.	Grade	Rec.	Pb	As	Hg
			(%)	(%)	(g/t)	(%)	(g/t)	(%)	(%)	(%)	(%)
FLC1	JL1	4.1	61.1	51	1.4	0.5	41	2.7	1.2	0.5	-
FLC3	1	2.7	50.8	82.5	5.2	1.5	250	15	8.9	1.4	205
FLC4A	3	9.3	58.7	84.1	1.1	0.6	41	5.3	1.1	0.5	175
FLC5	6	7.1	54.5	73.5	2.3	2.9	88	9.8	2.1	1.1	264
		5.8	56.3	72.8	2.6	1.4	105	8.1	3.3	0.9	215
		6.4	54.7	80.0	2.9	1.7	126	9.9	4.0	1.0	215
							Py-Au Rougher + Scavenger Concentrate
Test	Comp.	Wt.	Zinc		Gold		Silver			Grades
ID	ID	(%)	Grade	Rec.	Grade	Rec.	Grade	Rec.	Pb	As	S
			(%)	(%)	(g/t)	(%)	(g/t)	(%)	(%)	(%)	(%)
FLC1	JL1	62.4	3.3	42.4	15.5	90.0	62.9	62.6	1.9	13.5	24.1
FLC3	1	64.8	0.3	11.6	12.9	92.1	32.8	46.6	0.6	15.1	17.3
FLC4A	3	-	-	-	-	-	-	-	-	-	-
FLC5	6	48.5	0.4	4.5	8.7	74.2	11.8	10.0	0.4	10.9	14.5
		58.6	1.4	19.5	12.4	85.4	35.8	39.7	1.0	13.2	18.7
Average (all)Avg. FLC3, 4A, 5Average (all)Avg. FLC3, 5		56.7	0.4	8.1	10.8	83.2	22.3	28.3	0.5	13.0	15.9

Table 13.7 Summary of the Main Zone Variability LCT Results

Table 13.8 Locked Cycle Float Tests – Main Zone Calculated Heads

					Calculated Heads
TestID	Comp.ID	Wt.(%)	Zn(%)	Au(g/t)	Ag(g/t)	Pb(%)	As(%)	S(%)
FLC1	JL1	100.0	4.9	10.8	62.6	3.0	8.8	17.4
FLC3	1	100.0	1.6	9.1	45.3	2.0	10.0	12.6
FLC4A	3	100.0	6.5	16.2	71.3	5.0	9.8	17.6
FLC5	6	100.0	4.9	5.6	61.2	3.6	5.5	10.8
Average (all)		100.0	4.5	10.4	60.1	3.4	8.5	14.6
	Avg. FLC3, 4A, 5	100.0	4.3	10.3	59.3	3.6	8.4	13.7
	Grade Comp.1,3,6		5.7	10.8	86.8	3.8	8.1	13.9
M&I Resources			3.4	5.6	53.4	1.9

The total precious metal recovery into all three concentrate products was high, although highly variable between the various flotation products. Most of the remaining precious metals

following lead and zinc flotation were scavenged into the arsenopyrite / pyrite bulk concentrate, resulting in an overall 97% to 99% of the gold being floated. To ensure that the gold grade of the sulphide concentrate is higher than 15 g/t, which is the target grade above which a gold concentrate will be saleable, a cleaner circuit including a regrind mill is included in the PEA. The cleaner mass pull will be dictated by the sulphide content of the rougher/scavenger concentrate and based on the LCT work described above, will be approximately 50%, which will produce a product containing about 90% sulphides.

13.1.2.6 Flotation – Yellowjacket Zone

A series of open circuit flotation tests was undertaken that investigated a number of parameters including grind size, retention times, regrinding requirements, cleaning stages and reagents. Primary grind rougher tests using P80 grinds of between 36 and 74 microns showed no loss in performance with a coarser grind. A grind of 60-70 microns was selected for further studies although it was acknowledged that there may be an opportunity to increase the primary grind without adversely affecting Pb, Zn and Ag recoveries.

The results of the open circuit flotation cleaner test have been summarized in Table 13.9.

			Cleaner Grades					Rougher Recoveries
TestID	Description	Wt.	Ag	Pb	Zn	Wt.	Ag	Pb	Zn
		(%)	(g/t)	(%)	(%)	(%)	(%)	(%)	(%)
				Lead Circuit
F7	Composite 5	2.0	1,294	71.3	6.8	7.6	82.6	90.4	11.3
F9	Composite 3	2.6	794	38.9	2.7	11.7	88.9	90.1	12.3
				Zinc Circuit
F7	Composite 5	10.9	42.3	0.60	63.5	16.7	13.6	5.9	87.8
F9	Composite 3	5.4	37.1	0.56	64.6	10.6	7.1	4.3	85.7
				Final Tailing
F7	Composite 5	-	3.0	0.13	0.12	75.7	3.8	3.7	0.9
F9	Composite 3	-	4.0	0.12	0.12	77.7	6.0	5.6	1.9
			Composite Feed Assays
F7	Composite 5 –calculated head	100.0	60.4	2.71	9.52	-	-	-	-
	Composite 5 –measured head		62.1	2.68	7.96	-	-	-	-
F9	Composite 3 –calculated head	100.0	51.8	1.64	5.03	-	-	-	-
	Composite 3 –measured head		52.6	1.68	4.67	-	-	-	-

Table 13.9 Summary of Yellowjacket Open Circuit Flotation Cleaner Test Results

A single locked cycle flotation test (FLC1) was completed using the Yellowjacket Composite 5 sample. The test protocol generally followed the same base metal flotation procedure as for the Main Zone and used conditions developed during the open optimization circuit tests. The flowsheet used is presented in Figure 13.3 and results are summarized in Table 13.10.

Figure 13.3 Yellowjacket Zone Locked Cycle Flotation Test FLC1 Flowsheet

Source: F. Wright, March 2014, Appendix YJ-4, Locked Cycle Flotation Test FLC1.

Table 13.10
Summary of Flotation Test Results –YellowjacketComposite 5LCT1

				Grades				Metal Distribution (%)
Product	Wt.(%)	Au(g/t)	Ag(g/t)	Zn(%)	Pb(%)	As(ppm)	S(%)	Au	Ag	Zn	Pb
Pb con	4.1	1.57	1,090	10.5	43.3	740	12.1	58.3	81.9	5.4	87.5
Zn con*	12.0	0.25	56	61.9	1.1	91	32.3	27.5	12.3	93.1	6.7
Zn Sc. Tail	5.0	0.05	17	0.7	1.1	593	0.7	2.3	1.5	0.5	1.2
Ro. Tail	78.9	0.02	3	0.1	0.5	98	0.2	11.9	4.3	1.1	4.6
Feed (calc.)	100.0	0.11	55	8.0	0.1	149	4.6	100.0	100.0	100.0	100.0

*The mercury content of the zinc concentrate was 760 ppm.

13.1.2.7 Flotation Concentrate Quality

Examples of the analyses for the three flotation concentrates are provided in Table 13.11. The concentrates analysed were the final cycle from the two tests named FLC1, one that used Main Zone Composite JL1 and the one that used Yellowjacket Zone Composite DC7 & LG.

Element or	Units	ZincConcentrate		LeadConcentrate		GoldConcentrate
Compound		Main Zone	YJ Zone	Main Zone	YJ Zone	Main Zone
Zn	%	60.7	61.2	9.1	10.1	3.2
Pb	%	1.3	1.2	58.9	44.7	1.9
Au	g/t	1.4	0.3	30	1.3	16.7
Ag	g/t	41	122	730	1,111	67
Fe	%	6.5	1.7	5.8	1.0	21
Cu	%	0.21	0.13	1.8	0.6	0.05
S	%	32.5	32.0	18.5	12.5	21
As	%	0.5	0.01	3.0	0.07	11.7
Sb	%	0.04	0.01	1.3	0.9	0.12
Hg	ppm	290	720	45	-	22
F	g/t	200	-	400	-	3700
Cl	g/t	<1.8	-	<1.8		<1.8
SiO2	%	0.3	-	1.4		14

Table 13.11 Final Flotation Concentrate Multi-Analyses for Main Zone Test FLC1 and Yellowjacket FLC1

13.1.2.8 Acid Base Accounting

The float products from the heavy liquid tests (HLS5 to 10) were submitted for preliminary acid-base accounting tests. The Sobek acid consumption test was used to determine the negative net neutralization potential (NNP) of these samples. The NNP of the composites 1 to 5 varied between 136 to 935 kg CaCO3 equivalent while Composite 6 was negative 33 kg CaCO3. likely reflected by lower carbonate content as compared to Composites 1 to 5. The overall positive NNP of the samples suggests a low likelihood to generate acid rock drainage (ARD) from discharged HMS floats from the Main Zone.

Tailings

Preliminary characterization was performed on the final tailings generated from lock cycle flotation for both the Yellowjacket and Main Zones. This included both settling testwork and acid base accounting (ABA).

Sobek ABA tests performed on the Composite 3 Main Zone tailing gave an NNP of 123 kg CaCO3 equivalent. The NNP of the Main Zone master composite tailing averaged 117 kg CaCO3 equivalent. The NP to acid potential (AP) ratio ranged significantly from 3:1 to 20:1, primarily due to variation in the tailings sample sulphide content.

Acid base accounting (ABA) using the Sobek method on the Yellowjacket tailing gave a neutralization potential (NP) of 231 kg equivalent CaCO3, with a net neutralization potential (NNP) of 225 kg CaCO3 equivalent corresponding to a high NP:AP ratio of about 40:1. This indicates a low likelihood of acid rock generation potential, attributed to the high calcite content of the tailing.

Overall, the results from preliminary characterization tests on the tailings samples suggest limited potential technical and environmental implications for the storage of tailings.

13.1.2.9 Tailings Settling Tests

Standardized settling procedures were incorporated after initial beaker scoping evaluation using samples of final tailings generated from the locked cycle flotation tests for both the Yellowjacket and Main Zones.

Despite the relatively fine particle size of the Main Zone tailing (P80 ~40 microns), the material showed good settling characteristics and low observed turbidity in the supernatant after 24 hours. The calculated unit thickener area using the modified Coe and Clevenger Method was 0.1 m2 / t/d solids, with a solids settling rate of 32.2 m/d. Solids SG of the tailing was 2.74 and terminal pulp density calculated at 65% solids.

Yellowjacket tailing were coarser than the Main Zone (P80 ~70 microns) and showed better settling characteristics. Yellowjacket Composite 5 tailing exhibited a faster average settling rate with a calculated unit thickener area requirement of 0.05 m2 /t/d solids. There was also a lower supernatant turbidity of 15.5 mg/L and slightly higher terminal density of 67% after 24 hours.

13.1.3 Recent Flowsheet Development Testwork (2020)

A series of preliminary bench scale metallurgical tests were undertaken in 2020 at Base Met Labs under the direction Canenco using Main Zone composite JL-1. The objective of these tests was to investigate gold recovery by gravity separation and to improve the gold recovery into a high-grade gold sulphide concentrate using the concept developed by Bacon, Donaldson and Associates Ltd. in 1991. This concept comprised the production of an arsenopyrite concentrate which contained high gold values.

Test results documented in the November, 2020 Letter report from Base Met Labs indicated that a gravity concentrate containing 39 g/t Au, 100 g/t Ag and 23% As could be produced from flotation feed material with a gold recovery of about 27%.

Following the sequential flotation of lead (galena), pyrite and zinc (sphalerite), approximately 38% of the gold was recovered into the zinc tailings that comprised mainly arsenopyrite, and graded approximately 33 g/t Au, 25 g/t Ag and 38 % As.

Additional gravity tests were undertaken using a combined sample of the primary gravity concentrate and the zinc tailings. The gold grade / recovery model produced from this work is presented in Figure 13.4.

Figure 13.4 Gold Gravity Concentrate Grade vs Recovery Curve

The results from this preliminary gravity testwork show that a high-grade gold product can be produced albeit with an arsenic content of around 30% and a sharp drop in gold recovery to a concentrate containing above 40 g/t Au.

13.1.4 Recent Precious Metals Extraction Testwork (2011-2013)

Gold mineralization at the Main Zone is associated primarily with arsenopyrite and from historical work it has been shown to be highly refractory to direct cyanidation, with gold recoveries consistently below 25%. A confirmation cyanide leach test (C1) using ultrafine grinding to 80% passing less than 10 microns was performed on a gold concentrate assaying 16.8 g/t Au and 23.4 g/t Ag. After 48 hours of cyanidation there was a dissolution recovery of 18% for gold, and 83% for silver. The Yellowjacket Zone has low gold credits associated with base metal concentrates and was not evaluated for pre-treatment or cyanide leach recovery.

Historical testwork gave gold and silver extractions from arsenopyrite concentrates treated by pressure oxidation and cyanidation of approximately 90% and 50%, respectively. Extractions for arsenopyrite concentrates treated by bio-oxidation and cyanidation were approximately 90% and 80%, respectively.

The two principal cyanide pre-treatment steps considered in the recent phase of testwork for the Main Zone material comprised bio-oxidation (bioleach) and pressure oxidation (POX).

13.1.4.1 Bio-oxidation

Bio-oxidation focused on treatment of the gold enriched iron sulphide float concentrate. The gold concentrate did not respond well to float cleaning since the bulk concentrate consisted of massive arsenopyrite and pyrite and any rejection of sulphides resulted in gold losses. Several bioleach tests were performed at Inspectorate, as well as confirmation testing at SGS Laboratories in Lakefield Ontario.

The findings from both laboratories showed that gold cyanide recoveries following biooxidation could be improved significantly to the upper eighty percent to low ninety percent range for gold, depending on the test conditions and sample submitted. Silver cyanide recovery on the bio-residue was generally mid-nineties percent.

Negative issues were the long retention time and low pulp density required relative to typical bioleach conditions for arsenopyrite concentrate. Despite repeated attempts by multiple laboratories at adaptation of arsenopyrite cultures, the pulp density remained at 6 wt. % solids or less. Even using these relatively low solids densities, the batch bioleach tank retention times at Inspectorate achieved only 78% to 87% sulphide oxidation after 50 days while SGS achieved 95% sulphide oxidation but only after 69 days. Also, cyanide and lime requirements were high, possibly due to poor kinetics that can generate high cyanicides such as elemental sulphur in the bio-residue.

The poor response to bio-oxidation is postulated to be due to the high arsenic content coupled with other base metals and detrimental elements such as antimony that may be causing a synergistic negative effect on the microbes. Based on these preliminary bio-oxidation results, pressure oxidation became the primary focus.

13.1.4.2 Pressure Oxidation

Initial pressure oxidation (POX) tests were performed on iron sulphide flotation concentrates and some combined gold/zinc concentrates, produced from the various flotation tests (both open circuit and LCT) undertaken using Main Zone Composite JL1. The typical range of analyses of the material feed used for the POX testwork program were 17 to 25 g/t Au, 14% to 18% As, and 23% to 30% sulphide sulphur.

Standard base case conditions for autoclaving comprised:

Initial pH adjustment to 2.1 with sulphuric acid.
Temperature between 200 to 220 ̊C.
Pressure of around 415 psig (at 100 psig O2).
Retention time, base case 60 minutes.
Solids density, base case 10% by weight.

The first six POX tests (POX 1 to POX 6) investigated the retention time in the autoclave although two tests also considered 15% solids feed density. Tests POX 7 to POX 10 (four tests) examined the effect of "lime boil" to improve silver extraction and the final two tests (POX 11A and POX 11B) used zinc rougher scavenger tailing test samples as a feed source. A summary of all these test results is presented in Table 13.12.

Test ID	Description	Time	Temp	Solids	LBTime	(g/t)	Calc. Head		Consume(kg/t)		Extraction (%)
		(h)	(ºC)	(wt.%)	(min)	Au	Ag	NaCN	Lime	Au	Ag
POX1	Test FLC1 Au con	60	220	10	NA	17.5	137	35	56	98	16
POX2	Test FLC1 Au con	90	220	10	NA	19.5	147	20	41	96	34
POX3	Test F9 Zn+Au cons	60	220	10	NA	21.0	47	47	41	99	57
POX4	Test F9 Zn+Au cons	90	220	10	NA	17.2	39	55	47	95	17
POX5	Test F10 Au cons	15	220	15	NA	24.6	75	12	2	34	52
POX6	Test F10 Au cons	30	220	15	NA	25.1	67	15	2	63	62
POX7	As POX 1 + LB	60	220	20	290	20.6	92	7	72	89	93
POX8	Test FL3 Au con +LB	60	220	20	255	18.2	65	3	195	95	49
POX9	Test FL4 Au con +LB	60	220	20	185	26.1	28	2	424	87	63
POX10	Test FL5 Au con +LB	60	220	20	305	11.2	35	5	236	90	94
POX11A	Zn Ro.Sc Tail	60	220	20	NA	33.2	14	8	0.4	99	88
POX11B	Zn Ro.Sc Tail + LB	60	220	20		28.6	7	24	738	99	83

Table 13.12 Summary of POX Testwork Results Including Precious Metals Cyanide Leach Extractions

Notes: LB = Lime boil. Cyanide leaching for 48 h except POX9 which was 24 h.

Good gold recoveries (>95%) were achieved by pressure oxidation of the iron sulphide flotation concentrate followed by neutralization and cyanide leaching. Silver recoveries improved with a lime boil step prior to the cyanide leach although an economics of including this step need to be reviewed as, unlike the gold, more silver is recovered to the base metal concentrates and there appears to be a potential small gold loss with the introduction of this this process step, although this needs to be validated with more tests.

13.2 RECOMMENDATIONS

Additional metallurgical and mineralogical characterization studies are recommended in order to further develop and optimize the process flowsheet. Micon recommends the following:

Mineralogical and petrology investigations should be completed using composite samples that represent the various mineralization types and lithologies that occur within the Main and Yellowjacket deposits. Considering the nature of the mineralogy, Micon recommends that elemental deportment, mineral associations and mineral liberation by particle size be investigated. The information gleaned from this work will assist in guiding the future metallurgical testwork programs with the goal of optimizing recoveries of the various valuable components that occur within the deposit.
Sorting should be investigated as a replacement of the proposed HMS circuit. Upgrading using sorting technology will eliminate the challenges of operating an HMS system at a flotation concentrator. Although as yet there has been no testing to assess

the level of effectiveness of sorting on the Main Zone and Yellowjacket mineralization, there is potential that sorting, if reasonably successful, could positively impact the economics of the project.

Sequential flotation of zinc and lead from both the Main and Yellowjacket mineralization needs to be optimized. Particular efforts should be made to rationalize the flotation reagents selection and the reduction of usage rates.
Indications from independent concentrate marketing consultants suggest that the gold concentrate with a significant arsenic content needs to contain a minimum of 30 g/t Au to be saleable and therefore the bulk pyrite/arsenopyrite rougher/scavenger concentrate will need to be upgraded. Recent preliminary testwork using gravity separation technology and selective flotation to produce an arsenopyrite product showed that this gold grade can be readily achieved. Additional testwork looking at gravity separation, bulk sulphide/gold rougher concentrate cleaning and selective pyrite removal optimization, should be considered.
Further testwork to develop the oxidation of the gold concentrate and subsequent gold recovery is recommended. This includes additional development of pressure oxidation including the use of lower temperatures to reduce the cost of the autoclave, and possibly investigating alternative technologies as such as bio-oxidation.
Additional geochemical characterization studies on all tailings and residue streams from the selected flowsheet are recommended.
Thickening and filtration tests are recommended for all concentrate and residue streams where dewatering is required. This dewatering testwork on tailings and residue streams will be required to support the concept of dry-stack tailings deposition.
Once a mineral processing flowsheet has been finalized the metallurgical variability of the Main Zone deposit needs to be investigated.

14.0 MINERAL RESOURCE ESTIMATE

14.1 INTRODUCTION

The Mineral Resource Estimate presented herein is reported in accordance with the Canadian Securities Administrators' National Instrument 43-101 and has been estimated in conformity with generally accepted CIM "Estimation of Mineral Resource and Mineral Reserves Best Practices" guidelines. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no guarantee that all or any part of the Mineral Resource will be converted into Mineral Reserve. Confidence in the estimate of Inferred Mineral Resources is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure. Mineral Resources may be affected by further infill and exploration drilling that may result in increases or decreases in subsequent Mineral Resource Estimates.

All Mineral Resource estimation work reported herein was carried out by F. H. Brown, P.Geo., an independent Qualified Person as defined by National Instrument 43-101 by reason of education, affiliation with a professional association and past relevant work experience. The effective date of this Mineral Resource estimate is January 27, 2020. A draft copy of this Technical Report was reviewed by Rokmaster for factual errors.

Mineral Resource modeling and estimation were carried out using GEOVIA GEMS™ and Snowden Supervisor™ software programs.

14.2 DATA SUPPLIED

All sampling data were supplied as a Microsoft Access format database containing collar, survey, assay, bulk density and lithology data. A topographic surface and AutoCAD format wireframes of the underground workings were also supplied. All spatial data are reported relative to UTM NAD 83, Zone 11N.

As implemented by P&E, the database contains 582 unique collar records, encompassing surface trenches, underground chip sampling and drilling (Table 14.1). Of the 582 records, 29 records contained no associated assay data, were outside the Project limits, or were incomplete, and therefore were not used for Mineral Resource estimation.

Sample Type	Record Count	Total Metres
Drilling	298	41,135.22
Underground Chip Sampling	223	529.15
Surface Trench Sampling	32	85.57
Not used	29	140.41

Table 14.1 Database Summary

14.3 DATABASE VALIDATION

Industry standard validation checks were completed on the supplied databases. P&E typically validates a Mineral Resource database by checking for inconsistencies in naming conventions or analytical units, duplicate entries, interval, length or distance values less than or equal to zero, blank or zero-value assay results, out-of-sequence intervals, intervals or distances greater than the reported drill hole length, inappropriate collar locations, and missing interval and coordinate fields. Several minor out-of-sequence errors were detected and corrected. P&E independently acquired all 2010 and later assay results directly from the various laboratories and utilized that data for mineralized tenor validation of the pre-2010 assays. P&E believes that the supplied database is suitable for Mineral Resource estimation.

14.4 BULK DENSITY

The supplied database contains a total of 396 bulk density measurements. Representative samples of dry halved drill core from within the Main Zone and the margins of the Main Zone were selected for measurement. The dry weight of the drill core sample was weighed, and then the volume of displaced water determined from submerged drill core. Bulk density was calculated from the ratio of the dry weight of the drill hole core to the weight of the displaced water. The supplied bulk density measurements were used to estimate block model density values (Table 14.2).

In addition, P&E collected eighteen bulk density measurements from drill hole core for verification purposes, which are in agreement with values previously reported.

Statistic	Marginal	Main Zone	P&E
Count	396	256	18
Minimum	2.61	2.66	2.76
Maximum	5.03	5.00	4.18
Average	3.31	3.52	3.36

Table 14.2 Bulk Density Values

14.5 ECONOMIC PARAMETERS

Potentially economic mineralization in the January, 2020 block model was identified by calculating a net smelter return (NSR) value for each block based on individual assay values and the economic parameters listed in Table 14.3 and Table 14.4. NSR values were then used to construct economic mineralization domains. NSR values were calculated as:

Main Zone NSR = (Pb% x $21.16) + (Zn% x $22.01) + (Ag g/t x $0.52) + (Au g/t x $49.36) - $20.68 Yellowjacket Zone NSR = (Pb% x $19.58) + (Zn% x $22.93) + (Ag g/t x $0.48) + (Au g/t x $48.82) - $20.68

An NSR cut-off of CDN$110 per tonne was derived from $75/t mining, $25/t processing and $10/t G&A costs.

Table 14.3
Revel Ridge Economic Parameters –Main Zone

	Revel Ridge Project Main Zone - NSR Calculation					Jan 25/20
Dec 31 2019 24 mo Trailing Average Prices	Metal Price	Concentrate	Smelter	Refining Chg.	Refining Chg.	Average Grade
Element	$US/lb or oz	Recovery	Payable	$US/lb or oz	$C/lb or oz	% or aft
PЬ	$0,96	80%	95%	$0.00	$0.00	1.00%
Zn	$1.24	72%	85%	$0.00	$0.00	1.00%
Ag	$15.95	88%	91%	$0.50	$0.66	1.0
Au	$1.331	92%	96%	$10.00	$13.16	1.00
SC/SUS		50.760
Concentration Ratio (PB/Zn Blended)		20
Smelter Treatment Charge $US/dmt (Pb/Zn Blended Cost)		$185
Concentrate Shipping Charge $C/tonne		$65
Moisture Content		8%
U.S. NOON (ALC)	Payable Metal
Element	$C/tonne/g or %
PЬ	$21.16
Zn	$22.01
Au	50.52
Au	$49.36
	$93.06
Less Local Ore Haulage Cost to Mill	$5.00
Less Smelter Treatment Charges	$12.17
Less Concentrate Shipping Charges	$3.51
Penalties	$20.68

Table 14.4 Revel Ridge Economic Parameters – Yellowjacket Zone

	Revel Ridge Project Yellow Jacket Zone - NSR Calculation					Jan 25/20
Dec 31 2019 24 mo Trailing Average Prices	Metal Price	Concentrate	Smelter	Refining Chg.	Refining Chg.	Average Grade
Element	$US/lb or oz	Recovery	Payable	$US/lb or oz	$C/lb or oz	% or aft
PЬ	$0.96	74%	95%	$0.00	$0.00	1.00%
Zn	$1.24	75%	85%	$0.00	$0.00	1.00%
Ag	$15.95	80%	91%	$0.50	$0.66	1.0
Au	$1.331	91%	96%	$10.00	$13.16	1.00
SC/SUS		50.760
Concentration Ratio (PB/Zn Blended)		20
Smelter Treatment Charge $US/dmt {Pb/Zn Blended Cost)		$185
Concentrate Shipping Charge $C/tonne		$65
Moisture Content		8%
	Payable Metal
Element	$C/tonne/g or %
PЬ	$19.58
Zn	$22.93
Au	50.48
Au	548 82
	$91.80
Less Local Ore Haulage Cost to Process Plant	$5.00
Less Smelter Treatment Charges	$12.17
Less Concentrate Shipping Charges	$3.51
Penalties	$20.68

14.6 DOMAIN MODELING

The Main Zone, Hanging Wall, Footwall and Yellowjacket Zones have been defined by geologists along the primary structure, based on underground sampling, drilling, geological mapping and grade continuity. Based on the supplied interpretations, domain models were generated by P&E from successive polylines spaced every ten metres and oriented perpendicular to the trend of the mineralization. The outlines of the polylines were determined by an NSR value of approximately CDN$110/tonne based on the updated commodity prices, with demonstrated continuity along strike and down dip, and include low-grade material where necessary to maintain continuity between sections. All polyline vertices were snapped directly to drill hole assay intervals, in order to generate a true three-dimensional representation of the extent of the mineralization. Domain wireframes were then clipped above the topographic surface. The resulting domains were used for rock coding, statistical analysis and compositing limits (Figure 14.1).

Figure 14.1 Isometric Projection of Mineral Resource Domains (Looking West)

14.7 COMPOSITING

Assay sample lengths for within the defined zones range from 0.03 m to 6.00 m, with an average sample length of 0.65 m. For the assay lengths two modes are apparent, with a primary mode occurring at 0.50 m and a secondary mode at 1.00 m (Figure 14.2). Due to the narrow widths of the mineralized structures a compositing length of 0.50 m was therefore selected for use for Mineral Resource estimation.

Figure 14.2 Dotplot of Constrained Assay Lengths

Length-weighted composites were calculated within the Main, Footwall, Hanging Wall and Yellowjacket Zones domains. The compositing process started at the first point of intersection between the drill hole and the domain intersected and halted upon exit from the domain wireframe. The wireframes that represented the interpreted domains were also used to backtag a rock code field into the drill hole workspace. Assays and composites were assigned a domain rock code value based on the domain wireframe that the interval midpoint fell within. A nominal grade of 0.001 was used to populate a small number of un-sampled intervals. Composites that were less than one-half of the compositing interval in length were discarded so as to not introduce a short sample bias into the estimation process. The composite data were then exported to extraction files for grade estimation. Only assay values and underground channel samples were extracted for Mineral Resource estimation, and all trench samples were excluded.

14.8 EXPLORATORY DATA ANALYSIS

P&E generated summary statistics for the composite data (Table 14.5). The correlation between grade elements was also examined for the Main Zone, indicating a high degree of correlation between Ag and Pb, and a moderate degree of correlation between Ag and Zn and Pb and Zn (Table 14.6).

In addition, a comparison was made between underground chip sample values and drill hole assay values after compositing. The results indicate no significant bias between the two sample populations except for very low-grade Au values (Figure 14.3).

Ag	Mean(g/t)	St Dev	CV	Median(g/t)	Minimum(g/t)	Maximum(g/t)	Count
Main Zone	48.3	66.8	1.4	20.1	0.001	598.6	2935
Footwall	15.9	22.5	1.4	6.6	0.001	121.9	188
Hanging Wall	37.1	45.8	1.2	19.1	0.001	235.6	301
Yellowjacket	62.9	63.0	1.0	46.8	0.200	478.9	522
Au	Mean(g/t)	St Dev	CV	Median(g/t)	Minimum(g/t)	Maximum(g/t)	Count
Main Zone	4.73	7.13	1.51	2.26	0.001	157.19	2935
Footwall	2.44	7.02	2.88	0.98	0.001	89.34	188
Hanging Wall	0.78	1.89	2.43	0.10	0.001	13.57	301
Yellowjacket	0.08	0.12	1.61	0.01	0.001	1.02	522

Pb	Mean	St Dev	CV	Median	Minimum	Maximum	Count
Main Zone	(%)1.68	2.52	1.50	(%)0.57	(%)0.0008	(%)18.56	2935
Footwall	0.34	0.59	1.70	0.11	0.0006	3.27	188
Hanging Wall	1.55	2.10	1.35	0.79	0.0010	11.98	301
Yellowjacket	2.63	2.63	1.00	2.03	0.0100	23.40	522
Zn	Mean(%)	St Dev	CV	Median(%)	Minimum(%)	Maximum(%)	Count
Main Zone	3.13	4.45	1.42	0.98	0.0010	34.58	2935
Footwall	0.38	0.71	1.89	0.06	0.0010	4.31	188
Hanging Wall	3.91	5.30	1.36	1.49	0.0010	31.76	301

Table 14.5 Composite Summary Statistics

Note: St Dev = standard deviation, CV = coefficient of variation.

Table 14.6
Main Zone Composite Correlation Matrix

Element	Ag(g/t)	Au(g/t)	Pb(%)	Zn(%)
Ag	1	0.375	0.856	0.531
Au	0.375	1	0.287	0.145
Pb	0.856	0.287	1	0.621
Zn	0.531	0.145	0.621	1

Figure 14.3 Main Zone QQ Plots for Drill Hole vs. Chip Composites

14.9 TREATMENT OF EXTREME VALUES

The presence of high-grade outliers for the composite data was evaluated by reviewing probability plots of the composite sample populations (Figure 14.4). Capping thresholds were selected by disintegration of the upper tail of the composite sample distribution. Composite grades were capped to the selected threshold values prior to estimation (Table 14.7).

Figure 14.4 Capping Analysis Plots

Table 14.7 Capping Thresholds

Item	Au(g/t)	Ag(g/t)	Pb(%)	Zn(%)
Cap	50.00	460.00	18.00	34.00
Number Capped	6	4	4	1
Percent Capped	0.2%	0.1%	0.1%	0.0%
Mean	3.70	47.81	1.73	3.76

Item	Au(g/t)	Ag(g/t)	Pb(%)	Zn(%)
Capped Mean	3.65	47.75	1.73	3.76
Percent Contribution	4%	1%	1%	0%
Mean Above Cap	86.48	526.89	19.96	34.58
Percentile	1.00	1.00	1.00	1.00
Metal Loss	1.5%	0.1%	0.1%	0.0%
CoV	1.78	1.34	1.44	1.35
Capped CoV	1.61	1.33	1.43	1.35

14.10 CONTINUITY ANALYSIS

Domain-coded, composited sample data were used for continuity analysis. Strike orientations for the domains were modeled using the known geometry of the mineralization. Dip and dip plane orientations were modeled using orientations developed from variogram fans, which were assessed for geological reasonableness. Anisotropy was modeled with an average southeasterly strike and a north-easterly dip.

Based on the analysis of the resulting semi-variograms a strike distance of 60.0 m, a dip distance of 60.0 m, and a cross-dip distance of 20.0 m was selected as appropriate for Mineral Resource estimation. Continuity ellipses based on the observed ranges were then generated and used as the basis for estimation search ranges, distance calculations and Mineral Resource classification criteria.

14.11 BLOCK MODEL

A rotated block model was established across the Property with the block model limits selected so as to cover the extent of the mineralized domains and the block size reflecting the generally narrow widths of the mineralized zones and the drill hole spacing (Table 14.8). The block model consists of separate models for estimated grades, rock code, percent, density and classification attributes and a calculated NSR block grade. A volume percent block model was used to accurately represent the volume and tonnage that was contained within the constraining grade domains. As a result, the Mineral Resource boundaries were properly represented by the volume percent model's capacity to measure infinitely variable inclusion percentages. The volume of the historical underground workings was deemed insignificant and was not depleted from the model.

Dimension	Minimum	Maximum	Number ofBlocks	Block Size(m)
X	421,800	423,300	150	10
Y	5,680,100	5,683,100	300	10
Z	2,100	4,000	190	10
Rotation		-45°

Table 14.8 Block Model Setup

14.12 ESTIMATION AND CLASSIFICATION

Block density values were calculated using a single pass. Anisotropic inverse distance squared (ID2) linear weighting of between three and six bulk density values was used for the estimation of individual block bulk density values.

Anisotropic inverse distance squared linear weighting of capped composite values was used for the estimation of block grades, with the anisotropy defined by the axes of the search ellipse.

A three-pass series of expanding ellipsoids with varying minimum sample requirements were used for sample selection, grade estimation and classification. Composite data used during grade estimation were restricted to samples located in their respective domains. Individual block grades were then used to calculate an NSR block model.

During the first pass, five to six composites from three or more drill holes or underground channel samples within a search ellipsoid defined by 50% of the observed continuity ranges were required for estimation.

During the second pass, three to six composites from two or more drill holes or underground channel samples within a search ellipsoid defined by 100% of the observed continuity ranges were required for estimation.

During the third pass, three to six composites from one or more drill holes or underground channel samples were required. The search ellipsoid was expanded to ensure that all blocks within the defined domains were estimated. Search parameters are summarized in Table 14.9.

Pass	Search Radius	Minimum Number ofSamples	Maximum Number ofSamples
1st	30 m x 30 m x 10 m	5	6
2nd	60 m x 60 m x 20 m	3	6
3rd	240 m x 240 m x 80 m	3	6

Table 14.9 Search Parameters

Mineral Resources were classified in accordance with guidelines established by the Canadian Institute of Mining, Metallurgy and Petroleum:

Measured Mineral Resource: "A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of Modifying Factors to support detailed mine planning and final evaluation of the economic viability of the deposit. Geological evidence is derived from detailed and reliable exploration, sampling and testing and is sufficient to confirm geological and grade or quality continuity between points of observation."

Indicated Mineral Resource: "An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics are estimated with sufficient confidence to allow the application of Modifying Factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit. Geological evidence is derived from adequately detailed and reliable exploration, sampling and testing and is sufficient to assume geological and grade or quality continuity between points of observation."

Inferred Mineral Resource: "An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity."

Detailed summaries of underground chip sampling were supplied, and the information is of sufficient quality to justify the use of Measured Resources (personal communication Paul Cowley, P.Geo.). Based on the information supplied, P&E therefore considers that there is sufficient drilling and sampling information, and that this information is of a sufficient quality, to support a Measured, Indicated and Inferred classification for the Revel Ridge Mineral Resource Estimate.

Mineral Resource classification was conducted by generating three-dimensional envelopes around those parts of the block model for which the drill hole data and grade estimates met certain criteria. The resulting classifications were iteratively refined to be geologically reasonable in order to prevent the generation of small, discontinuous areas of a higher confidence classification being separated by a larger area of a lower confidence Mineral Resources.

Measured Mineral Resources were defined based on the results of the first pass, and then consolidated into an envelope digitized around the central area of blocks estimated during the first pass. This process downgraded isolated higher confidence blocks and combined the Measured Mineral Resources into a continuous unit.

Indicated Mineral Resources were defined based on the results of the second pass, and then consolidated into an envelope digitized around the central area of blocks estimated during the second pass. This process downgraded isolated higher confidence blocks and combined the Indicated Mineral Resources into a continuous unit.

All remaining Main Zone blocks estimated were classified as Inferred, including all blocks in the Footwall Zone (Figure 14.5).

Figure 14.5 Isometric Projection of Block Classification (Looking West)

14.13 MINERAL RESOURCE ESTIMATE

The Updated Mineral Resource Estimate for the Revel Ridge Deposit is reported at an NSR cut-off value of CDN$110/tonne, with an effective date of January 27, 2020. The updated Mineral Resource Estimate reports 5.27M Measured and Indicated tonnes containing 1.28M gold-equivalent ounces and 4.96M Inferred tonnes containing 1.01M gold-equivalent ounces (Table 14.10).

Mineralized	Classification	Tonnes	Au	Au	Ag	Ag	Pb	Zn	Au Eq	Au Eq
Zone		(000't)	(g/t)	(koz)	(g/t)	(koz)	(%)	(%)	(g/t)	(koz)
Main Zone	Measured	1,352	6.13	266	62.8	2,730	2.19	4.09	9.14	397
	Indicated	2,848	5.33	488	49	4,487	1.72	3.11	7.56	692
	Meas & Ind	4,200	5.59	755	53.4	7,216	1.87	3.43	8.07	1,089
	Inferred	4,562	4.36	639	61.8	9,064	1.88	2.59	6.55	961
	Indicated	298	0.91	9	55.3	530	2.50	5.72	4.70	45
HW Zone	Inferred	38	0.22	0	75	92	3.08	5.44	4.34	5
FW Zone	Inferred	341	3.91	43	25.3	277	0.53	0.48	4.20	46
Yellowjacket	Indicated	771	0.09	2	62.6	1,552	2.60	9.93	NA	NA
Zone	Inferred	23	0.11	0	55.4	41	2.65	7.68	NA	NA

Table 14.10 Revel Ridge 2020 Mineral Resource Estimate (1-7)

Note: koz = thousands of ounces.

1. Mineral Resources which are not Mineral Reserves do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues.
1. The Inferred Mineral Resource in this estimate has a lower level of confidence than that applied to an Indicated Mineral Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of the Inferred Mineral Resource could be upgraded to an Indicated Mineral Resource with continued exploration.
1. The Mineral Resources in this estimate were calculated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.
1. The following parameters were used to derive the NSR block model cut-off values used to define the Mineral Resource: Dec 31, 2019 US$ two-year trailing avg. metal prices:
- Pb $0.96/lb, Zn $1.24/lb, Au $1,331/oz, Ag $15.95/oz
- Exchange rate of US$0.76 = CDN $1.00
- Process recoveries of Pb 74%, Zn 75%, Au 91%, Ag 80%
- Smelter payables of Pb 95%, Zn 85%, Au 96%, Ag 91%
- Refining charges of Au US$10/oz, Ag US$0.50/oz
- Concentrate freight charges of CDN$65/t and Smelter treatment charge of US185/t
- Mass pull of 5% and 8% concentrate moisture content.
1. NSR cut-off of CDN$110 per tonne was derived from $75/t mining, $25/t processing, $10/t G&A.
1. AuEq= Au g/t + (Ag g/t x 0.011) + (Pb % x 0.422) + (Zn % x 0.455)
1. Above parameters derived from 2012 PEA and other similar benchmarked projects.

P&E is of the opinion that the current Mineral Resource Estimate meets the reasonable prospect of economic extraction due to the approximate 7.0 g/t AuEq average grade and the $110/t NSR cut-off (equal to approx. 3.5 g/t AuEq). P&E has experience with other similar projects and is of the opinion that the cut-off grade and cost assumptions are reasonable.

P&E is not aware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors which may materially affect the Mineral Resource estimate. A material drop in metal prices below the Dec 31, 2019 two year trailing average prices used for the current Mineral Resource Estimate or a significant increase in operating costs could materially affect the cut-off and average grades and potentially result in a revised lower Mineral Resource Estimate tonnage

14.14 VALIDATION

The block model was validated visually by the inspection of successive section lines in order to confirm that the block model correctly reflects the distribution of high-grade and low-grade samples (Appendix). As a further check on the model the average model block grades were compared to a Nearest Neighbour (NN) model and the uncapped mean of the composite data (Table 14.11). No significant bias between the block model and the input data was noted.

Element	Model Mean	Uncapped CompositeMean	NN Mean
Ag g/t	44.73	48.3	44.06
Au g/t	4.60	4.73	4.46
Pb %	1.58	1.68	1.57
Zn %	2.88	3.13	2.86

Table 14.11 Main Zone Validation Statistics

In addition, local trends were evaluated by comparing the ID2 block estimates to a Nearest Neighbour estimate (NN) at zero cut-off along the strike of the Main Zone for Measured and Indicated Mineral Resource blocks (Figure 14.6). In general, the ID2 block estimates are in good agreement with the NN estimates and demonstrate no evidence of systematic bias in the model.

Figure 14.7 comprises a drill hole plan and five vertical cross-sections all showing the Main, Hanging Wall and Yellowjacket zones; the cross sections display the zone block models.

Figure 14.6 Main Zone Swath Plots

Figure 14.7 Drill Hole Plan and Vertical Cross-Sections

15.0 MINERAL RESERVE ESTIMATE

There are no mineral reserves at the Revel Ridge property.

16.0 MINING METHODS

16.1 SUMMARY

The potentially mineable portion of the Revel Ridge Yellowjacket and Main Zone mineral resources extends along the strike length of the deposit for over 1 km, and over 800 m vertically from the 1310 Level to the 510 Level.

The mineral resource is subdivided into the Main Zone and the Yellowjacket Zone. The mineable stope shapes for these two zones are shown along with the topography intersection at the 830 m level, the existing development and the overall block model in Figure 16.1 and Figure 16.2.

The basis for the mine plan is a block model developed by P&E (2020). The stopes were generated using Datamine's MSO software and applied the calculated stope cut-off grades against the NSR attribute included in the P&E block model.

Figure 16.3 depicts the mineable stope shapes for the Main Zone colour coded by their NSR value as per the P&E 2020 block model.

Figure 16.4 depicts the mineable stope shapes for the Yellowjacket Zone colour coded by their NSR value as per the P&E 2020 block model.

Figure 16.1 2020 Block Model and Stope Shapes: Plan View

Figure 16.2 2020 Block Model and Stope Shapes: Section View

Figure 16.3 Main Zone Stopes Included in the Mine Plan Coded for NSR Values

Figure 16.4 Yellowjacket Zone Stopes Included in the Mine Plan Coded for NSR Values

Currently, there is an exploration drift on the 830 Level, accessed by the 832 adit portal. There are also several other adits, none of which are anticipated to be used for mine traffic during mining operations but can be reopened and reinforced to provide secondary egress as required. Portions of the mine plan, above and below the 830 Level, are referred to as the upper and lower mine, respectively.

Figure 16.5 (over) illustrates the current condition of the portal which has recently been described by BGC Engineering (June 23, 2020) as being in generally good condition from a geotechnical standpoint.

Micon has developed a conceptual mine plan to estimate a resource extraction schedule over the twelve (12) year life-of-mine (LOM).

The mine design is based on stope shapes that exceed a net smelter return (NSR) cut-off value of $107/t for the Yellowjacket Zone and $133/t for the Main Zone.

These cut-off grades are based on an approximated $75/t mining cost for both zones, a $10/t G&A cost for both zones and a $22/t processing cost for Yellowjacket Zone mill feed and $48/t processing cost for the Main Zone mill feed.

Figure 16.5 Photo of 832 m Elevation Portal in September, 2020

The stope shapes meet these cutoff grades (CoG) inclusive of internal dilution within the stope in addition to an allowance for an overall average overbreak of 0.25m for the entirety of all hanging wall surface area.

The stope shapes used were 20 m in height, 20 m in length and had a minimum stope design width of 2.5 m. The stope shapes also include an allowance for overbreak and include internal waste dilution.

Figure 16.6 shows a schematic of Transverse Longhole Mining Method.

Source: Rokmaster.

Figure 16.6 Schematic Cross- and Longitudinal Section of Mining Method

Source:Rokmaster

The contents of the stope shapes for the Main Zone that are include in the mine schedule are presented in Table 16.1.

Zone	Code	Class	Tonnes	NSRavg	Ag(g/t)	Au(g/t)	Pb(%)	Zn(%)	Density
Main	100	Measured	1,597,883	365	51.32	4.98	1.77	3.26	3.53
Main	100	Indicated	3,156,425	310	40.26	4.45	1.42	2.57	3.65
Main	100	Inferred	4,476,358	293	56.13	3.89	1.72	2.39	3.58
Main -FW	200	Measured	0	0	0.00	0.00	0.00	0.00	0.00
Main -FW	200	Indicated	0	0	0.00	0.00	0.00	0.00	0.00
Main -FW	200	Inferred	81,881	218	24.34	4.12	0.54	0.39	3.43
Main - HW	300	Measured	0	0	0.00	0.00	0.00	0.00	0.00
Main - HW	300	Indicated	72,758	266	67.83	0.21	3.20	7.67	3.46
Main - HW	300	Inferred	7,567	210	66.63	0.18	2.73	5.66	3.46
Main	Total	Measured	1,597,883	365	51.32	4.98	1.77	3.26	3.53
Main	Total	Indicated	3,229,183	309	40.88	4.35	1.46	2.68	3.65
Main	Total	Inferred	4,565,806	291	55.58	3.89	1.70	2.36	3.58

Table 16.1 Stope Shape Tonnes and Grades: Main Zone

Note: these stope shape tonnes and grades include internal and external dilution from blocks external to the resource model.

The contents of the stope shapes for the Yellowjacket Zone that are include in the mine schedule are presented in Table 16.2.

Zone	Code	Class	Tonnes	NSRavg	Ag(g/t)	Au(g/t)	Pb(%)	Zn(%)	Density
YJ	420	Measured	0	0	0.00	0.00	0.00	0.00	0.00
YJ	420	Indicated	21,426	187	30.77	0.07	1.47	6.51	3.56
YJ	420	Inferred	0	0	0.00	0.00	0.00	0.00	0.00
YJ	460	Measured	0	0	0.00	0.00	0.00	0.00	0.00
YJ	460	Indicated	165,444	203	34.57	0.06	1.64	7.08	3.53
YJ	460	Inferred	0	0	0.00	0.00	0.00	0.00	0.00
YJ	470	Measured	0	0	0.00	0.00	0.00	0.00	0.00
YJ	470	Indicated	447,093	223	46.37	0.06	2.01	7.61	3.47
YJ	470	Inferred	0	0	0.00	0.00	0.00	0.00	0.00
YJ	480	Measured	0	0	0.00	0.00	0.00	0.00	0.00
YJ	480	Indicated	12,550	276	55.20	0.01	2.44	9.30	3.52
YJ	480	Inferred	0	0	0.00	0.00	0.00	0.00	0.00
YJ	Total	Measured	0	0.00	0.00	0.00	0.00	0.00	0.00
YJ	Total	Indicated	646,512	218	43.00	0.06	1.90	7.47	3.49
YJ	Total	Inferred	0	0.00	0.00	0.00	0.00	0.00	0.00

Table 16.2 Stope Shape Tonnes and Grades: Yellowjacket Zone

Note: these stope shape tonnes and grades include internal and external dilution from blocks external to the resource model.

NSR was included as a variable in the P&E block model and was used as the basis upon which the CoG was applied. The mine plan involves mechanized, long-hole stoping with truck haulage, and backfill using cemented process tailings and development waste.

Sills developed through orezones will be shaped in order to minimize waste using what are known as "Shanty Backs".

The production schedule was generated by Micon with the aid of Datamine's MSO software to generate stope shapes and followed a conceptual development plan that targeted levels requiring minimum development first and then individual stopes on those levels that were highest in Au grade first.

16.2 MINE PLAN

The conceptual production mine plan produces an average 2,300 t/d of process feed for almost twelve years (11.96 years).

The first nine (9) years of the project focusses solely on mining the Main Zone, while years ten through twelve incorporate the mining of Yellowjacket Zone as well. The mining rate of 2,300 tonnes per day amounts to 839,500 tonnes per year, for a total of 10,039,384 tonnes over the 12-year mine life. The process feed from the Main Zone amounts to 9,392,872 tonnes, while the remaining 646,512 tonnes of process feed comes from Yellowjacket Zone.

The mine plan mines on 40 different levels from the 510 m level to the 1310 m level which extends up to the 1330 m elevation. The level spacing is 20 m vertically.

The crosscuts that connect the drifts to the sills and act as draw-points are generally spaced at 40 m (with some exceptions where single 20 m long stopes are found).

Mill-feed from sills is included as a mining cost rather than development cost. Waste from sills is counted as an operational development cost as are crosscuts and drifts.

In year zero (0) a total ramp length of 667 m will be developed giving access down to level 780 and up to level 880 from the main development level on 830. The level accesses, along with some drifts and cross cuts will be started on levels 780, 800, 820 and 840 in order to facilitate the mining of sills and stopes in year one.

In year one (1), 839,500 tonnes will be mined from stope levels 780, 800, 820, 840, 860, and 880. The NSR per tonne mined for this period is CDN$412/t with a gold grade of 6.48 g/t, a silver grade of 47.28 g/t, a lead grade of 1.49 % and a zinc grade of 2.43 %. An additional 667 m will be developed in year one (1) to give access to the levels that will be mined in year two i.e., down to 740m level and up to the 940 m level.

In year two (2), 839,500 tonnes will be mined from stope levels 740, 760, 880, 900, 920, and 940. The NSR per tonne mined for this period is CDN$417/t with a gold grade of 6.10 g/t, a silver grade of 52.22 g/t, a lead grade of 1.82% and a zinc grade of 3.05%. An additional 400 metres of ramp will be developed down to the 700m level and up to the 960m level.

In year three (3), 839,500 tonnes will be mined from stope levels 700, 720, 940, and 960. The NSR per tonne mined for this period is CDN$433/t with a gold grade of 6.10 g/t, a silver grade of 61.49 g/t, a lead grade of 2.09% and a zinc grade 3.31%. An additional 667 m ramp will be developed down to the 640 m level and up to the 1000 m level.

In year four (4), 839,500 tonnes will be mined from stope levels 640, 660, 680, 960, 980, and 1000. The NSR per tonne mined for this period is CDN$421/t with a gold grade of 5.90, a silver grade of 62.75, a lead grade of 2.15% and a zinc grade of 3.15%. An additional 800 m ramp will be developed down to the 620 m level and up to the 1060 m level.

In year five (5), 839,500 tonnes will be mined from stope levels 780, 600, 620, 1000, 1020, 1040, and 1060. The NSR per tonne mined for this period is CDN$385/t with a gold grade of 5.49, a silver grade of 56.55, a lead grade of 1.80 and a zinc grade 2.90%. An additional 400 m ramp will be developed down to the 560 m level and up to the 1100 m level.

In year six (6), 839,500 tonnes will be mined from stope levels 560, 1040, 1060, 1080, and 1100. The NSR per tonne mined for this period is CDN$264/t with a gold grade of 3.46 g/t, a silver grade of 40.57g/t, a lead grade of 1.45% and a zinc grade 2.70%. There is 333 m of ramp development in year six.

In year seven (7), 839,500 tonnes will be mined from stope levels 640, 920, 940, and 980. The NSR per tonne mined for this period is CDN$228/t with a gold grade of 3.01 g/t, a silver grade of 34.40 g/t, a lead grade of 1.27% and a zinc grade of 2.33%. There is 333 m of ramp development in year seven.

In year eight (8), 839,500 tonnes will be mined from stope levels 600, 620, 640, 1020, 1040, 1060 and 1080. The NSR per tonne mined for this period is CDN$241/t with a gold grade of 2.95 g/t, a silver grade of 58.88 g/t, a lead grade of 1.55% and a zinc grade of 2.26%. There is 333 m of ramp development in year eight.

In year nine (9), 839,500 tonnes will occur in in the main zone. Levels mined include the 560, 580, 660, 680, 760, 1000, 1080 and 1100. The NSR per tonne mined for this period is CDN$229/t with a gold grade of 2.67 g/t, a silver grade of 52.58 g/t, a lead grade of 1.58% and a zinc grade of 2.42%. An additional 467 m ramp will be developed.

In year ten (10) mining will occur in both Main Zone (677,901 tonnes) and Yellowjacket Zone (161,598 tonnes). A combined total of 839,500 tonnes will be mined from stope levels 540, 720, 760, 800,820, 860, 960, 1100, 1120, 1140, 1160, 1180, 1220, 1240, 1280, and 1300. The NSR per tonne mined for this period is CDN$217/t with a gold grade of 2.41 g/t, a silver grade of 39.36 g/t, a lead grade of 1.36% and a zinc grade 3.02%. An additional 133 m ramp will be developed.

In year 11 mining will occur in both the Main Zone (679,021 tonnes) and Yellowjacket Zone (160,479 tonnes). A combined total of 839,500 tonnes will be mined from stope levels 500, 520, 540, 700, 740, 780, 800, 820, 840, 880, 1200, and 1260. The NSR per tonne mined for this period is CDN$216/t with a gold grade of 2.07 g/t, a silver grade of 37.89 g/t, a lead grade of 1.43%and a zinc grade 3.59%.

In year 12 mining will occur in both the Main Zone (480,449 tonnes) and Yellowjacket Zone (324,432 tonnes). A combined total of 804,884 tonnes will be mined from stope levels 660, 700, 720, 740, 760, 780, 800, 820, 860, 880, 1120, 1140, 1160, 1180, 1220, and 1240. The NSR per tonne mined for this period is CDN$180/t with a gold grade of 0.85 g/t, a silver grade of 48.35 g/t, a lead grade of 1.76% and a zinc grade 4.12%.

The mining rate assumed throughout the mine life will be stable at 2,300 t/d.

For the first nine years all the production mill feed will come from the Main Zone.

In year ten when mining begins in the Yellowjacket Zone, approximately 678 kt will come from main zone while 162 kt will come from Yellowjacket Zone.

In year eleven 679 kt will come from main zone while 160 kt will come from Yellowjacket Zone.

Finally, in year twelve 480 kt will come from main zone while 324 kt will come from Yellowjacket Zone.

The mine plan initially targets levels that require the minimum amount of ramp and vertical raise development.

The air raises for the entire mine will be raise bored using contractors during the first four years of the mine life after which all raises will be 3 m x 3 m traditional raises. Adequate fresh air raises will be developed during year zero in order to facilitate mining in year one. Ore-passes will also be raise-bored during the first 4 years of the mine life.

The development schedule (Table 16.3) illustrates the plan for ramp and level development for mining of the different levels on an annual basis.

StopeLevel	Y0	Y1	Y2	Y3	Y4	Y5	Y6	Y7	Y8	Y9	Y10	Y11	Y12
1300										R	M
1280										R	M
1260										R		M
1240									R		M		M
1220									R		M		M
1200								R				M
1180								R			M		M
1160							R				M		M
1140							R				M		M
1120						R					M		M
1100						R	M			M	M
1080						R	M		M	M
1060					R	M	M		M
1040					R	M	M		M
1020					R	M			M
1000				R	M	M				M
980				R	M			M
960			R	M	M						M
940		R	M	M				M
920		R	M					M
900		R	M				M
880	R	M	M				M					M	M
860	R	M					M				M		M
840	R	M					M					M
820	R	M					M				M	M	M
800	R	M									M	M	M
780	R	M										M	M
760		R	M				M	M		M	M		M
740		R	M					M				M	M
720			R	M				M			M		M
700			R	M								M	M

Table 16.3 Development Plan for Stope Levels

StopeLevel	Y3	Y4	Y5	Y6	Y7	Y8	Y9	Y10	Y11	Y12
680	R	M					M
660	R	M					M			M
640	R	M			M	M
620		R	M			M
600		R	M			M
580		R	M				M
560			R	M			M
540				R				M	M
520								R	M
500								R	M

Upper and lower levels will be developed simultaneously in order to give access to a sufficient number of levels and development faces so that there will always be three to five stopes in production on any given day. On average, four stopes should be in production with each producing approximately 575 t/d on average. These estimates should be validated through a detailed production study during the next phase of study.

Production mining above the 830 m elevation will use ore-passes to bring mill feed down to the main haulage level on the 830 m level by gravity, where it will exit the mine via the 832 m portal.

Backfill will primarily be cemented, however where appropriate the cost savings afforded by uncemented backfill will be used.

16.3 DEVELOPMENT LAYOUT

The development metres by level and type are estimated in Table 16.4.

Level accesses: 30 metres of level access are planned on each level from the ramp to the drift.

Drifts: drifts will run the length of the mineralization on each level from the northernmost to southernmost stope.

Stopes: the number of stopes on each level will determine the amount of cross-cut development.

Crosscuts: Crosscuts are assumed to be maintained at 40 m spacing in general although some will be at spacing dictated by the location of the stopes. No two draw points will be closer than 20m apart.

Safety Bays: three metres of safety bays are planned for every 50 metres of drift.

Remucks: ten metres of remuck area are planned on every level.

Sumps: Ten metres of sumps are planned on every fourth level above the 830 m elevation and on every level.

Workshops & Storage: Twenty metres of workshops and storage areas are planned on every fifth level.

Sills: sills will run the length of every stope intersecting with crosscut for a net length of 15 m per stope.

Level	LevelAccess	Drift	Stopes	CrossCut	Remucks	SafetyBays	Sumps	Workshop/ Storage	TotalwithoutSills	Sills	Total
elevation	m	m	#	m	m	m	m	m	m	m	m
1,310	30	50	3	23	10	6	10	0	129	15	144
1,290	30	100	5	38	10	6	0	20	204	30	234
1,270	30	320	11	83	10	19	0	0	462	45	507
1,250	30	320	16	120	10	19	0	0	499	120	619
1,230	30	400	19	143	10	24	10	0	617	75	692
1,210	30	400	18	135	10	24	0	0	599	75	674
1,190	30	800	21	158	10	48	0	20	1,066	135	1,201
1,170	30	800	27	203	10	48	0	0	1,091	165	1,256
1,150	30	1,010	32	240	10	61	10	0	1,361	210	1,571
1,130	30	1,035	37	278	10	62	0	0	1,415	255	1,670
1,110	30	1,100	40	300	10	66	0	0	1,506	300	1,806
1,090	30	1,180	42	315	10	71	0	20	1,626	330	1,956
1,070	30	1,100	44	330	10	66	10	0	1,546	330	1,876
1,050	30	1,090	48	360	10	65	0	0	1,555	360	1,915
1,030	30	1,110	53	398	10	67	0	0	1,615	390	2,005
1,010	30	1,190	57	428	10	71	0	0	1,729	405	2,134
990	30	1,250	58	435	10	75	10	20	1,830	435	2,265
970	30	1,290	63	473	10	77	0	0	1,880	480	2,360
950	30	1,300	63	473	10	78	0	0	1,891	450	2,341
930	30	1,210	60	450	10	73	0	0	1,773	375	2,148
910	30	1,250	51	383	10	75	10	0	1,758	375	2,133
890	30	1,430	52	390	10	86	0	20	1,966	375	2,341
870	30	1,570	50	375	10	94	0	0	2,079	365	2,444
850	30	1,570	53	398	10	94	0	0	2,102	325	2,427
830	0	0	50	375	10	6	10	0	401	310	711
810	30	1,410	61	458	10	85	10	0	2,003	350	2,353
790	30	1,430	65	488	10	86	10	20	2,074	325	2,399
770	30	1,430	61	458	10	86	10	0	2,024	405	2,429
750	30	1,300	66	495	10	78	10	0	1,923	375	2,298
730	30	1,180	61	458	10	71	10	0	1,759	400	2,159
710	30	1,180	58	435	10	71	10	0	1,736	390	2,126
690	30	1,180	54	405	10	71	10	20	1,726	350	2,076
670	30	1,140	50	375	10	68	10	0	1,633	370	2,003
650	30	1,000	49	368	10	60	10	0	1,478	345	1,823
630	30	1,000	44	330	10	60	10	0	1,440	300	1,740
610	30	940	33	248	10	56	10	0	1,294	225	1,519
590	30	820	30	225	10	49	10	20	1,164	195	1,359

Table 16.4 Development Metres by Level and Type

Level	LevelAccess	Drift	Stopes	CrossCut	Remucks	SafetyBays	Sumps	Workshop/ Storage	TotalwithoutSills	Sills	Total
570	30	600	29	218	10	36	10	0	904	210	1,114
550	30	600	14	105	10	36	10	0	791	90	881
530	30	340	7	53	10	20	10	0	463	45	508
510	30	13	1	8	10	1	10	0	72	15	87
Total	1,200	38,438	1,656	12,420	410	2,315	230	160	55,172	11,120	66,292

During the mine's twelve-year life, a total of:

1. 493 kt of waste will be produced from 5.3 km of 5 m x 5 m Ramps.
1. 116 kt of waste will be produced from 2,000 m of 3 m diameter bored raises and 2,000 m of traditional 3 m x 3 m raises.
1. 2,858 kt of waste will be produced from 55.2 km of 3.5 m x 4.0 m horizontal development.
1. From 11.1 km length of 5 m x 5 m sills 2,845 kt will be mined of which 2,299 kt will be sent as mill feed while the remaining 546 kt will be waste.

Initial waste development involves accessing both the upper and lower mine with ramps from the existing workings, as well as from a new portal.

An incline, at a grade of 15%, will provide access to the upper mine from the existing 832 level. The incline ramp for the upper section of the mine will ultimately extend to the 1310 Level (1330m elevation at top of the stope level).

A decline ramp at a grade of -15% %, will provide access to the lower mine from the existing 832. The decline ramp for the lower section of the mine will ultimately extend to the 510 Level.

Vertical ventilation raises will be added in order to ensure adequate air circulation to the areas of the mine furthest from the adits and in order to provide additional egress to points higher up on the terrain (in case lower egress points are blocked by avalanche or other causes).

From the main 832 level, a decline ramp will be driven with a gradient of -15% to access the lower mine.

Development in the lower levels in the mine will include sumps for dewatering pumps,

Development in the upper levels will include strategically placed ore-passes to exploit the potential of using gravity to help maximize production rates while minimizing haulage costs.

Figure 16.7 depicts the conceptual layout for lower mine development. The view is from the south looking toward the north. Existing workings are shown in pink, proposed development in blue, and the mineralized zone in brown.

Figure 16.7 Schematic Showing Conceptual Plan for Waste Development in the Lower Mine (Not to Scale)

Access to the mine will be through the existing 832 adit portal which will be used for haulage, as well as transport of workers and materials.

Mill-feed transport from the upper mine will use ore-passes in order to streamline mine traffic and achieve planned production.

Planned development metres and tonnes are presented in the schedule in Table 16.11.

Accordingly, in year zero (0) a full year of development is planned together with sill development that will facilitate a steady state of production and two months of stope inventory from the beginning of year one.

In subsequent years, a steady output at full rate is scheduled.

Over the life-of-mine, approximately 55,172 m of waste development is planned, including ramps, drifts, and raises.

A total of 11,120 metres of sills are planned.

16.4 STOPING

Micon has designed the Main Zone and Yellowjacket stopes using the 2020 block model provided by P&E and Datamine's MSO stope optimization software.

An NSR cut-off value of CDN$107/t was used for the Yellowjacket Zone based upon PEA level assumptions that the mining cost will average CDN$75/t mill feed, processing cost will average $22/t mill feed and G&A will average $10/t mill feed.

An NSR cut-off value of CDN$133/t was used for the Main Zone based upon PEA level assumptions that the mining cost will average CDN$75/t mill feed, processing cost will average $48/t mill feed and G&A will average $10/t mill feed.

Stope orientations output by the software were constrained by the geologic wireframe provided by P&E.

Additional input parameters for the stope optimizer were based upon the assertion that stope heights are confined to 20-m sublevels, with 5-m high sill development on each level.

Stope widths were allowed to follow the width of the vein to a minimum of 2.5 m. Dilution is accounted for in the stope optimization process, and varies with the width of the vein.

Longitudinally oriented stopes are planned, with cemented tailings and development waste as backfill. Mine sublevels, where drilling, blasting, and mucking will take place, are vertically spaced 20 m apart, with main haulage levels at 60 m vertical intervals. Sublevels are conservatively spaced to allow for gravity flow of broken material, given the intermediate dip of the mineralization (about 55o , or less). Stopes are also designed to aid in gravity flow. On the haulage levels, crosscuts will be driven every 40 m along strike to access the muck.

To the extent possible, mining will progress in an overhand fashion (drilling uppers only) with the lower blocks being mined first and progressively higher blocks being mined subsequently. The stoping cycle will include long-hole drilling, blasting, mucking, and backfilling.

Backfill plant construction will commence at the beginning of the first year, most likely on the 830 level. Both development waste and cemented backfill will be used. Floats from the HMS plant will also be available for use in backfill.

PAG waste will report directly to the PAG Waste Pile and will not be used for backfilling in the upper mine. Where the schedule permits, some PAG waste may be stowed in areas of the lower mine that will be permanently flooded following mine closure.

Combined sill development and stoping are expected to provide 2,300 t/d of mill feed. Over the life-of-mine, sill development will contribute about 2.30 million tonnes and stoping will contribute about 7.74 million tonnes for a combined total of 10.04 million tonnes.

Stopes are expected to remain stable between mining and backfilling. Geomechanical evaluation by KCBL in 2007 (Klohn Crippen Berger Limited, 2007) was based on the CSIR Classification System:

"The overall combined rating of the rock mass was 65/100, which is Class II rock and indicative of 'good rock.' The average stand-up time, for unsupported rock with a 4 m span is 6 months and the rock mass has high cohesion (>150 kPa) and a friction angle (40-45 degrees), again indicative of good quality rock."

Further geotechnical test work will be undertaken to determine the levels of ground support needed.

16.4.1 Main Zone

Mining dilution in the Main Zone comprises an average 6.1 % allowance for overbreak and a further 26.1% for internal (waste or marginal grade mineralization) included within the stope shapes where operationally necessary, for a total dilution of 32.1% included within the mine schedule.

In the main zone a total of 1579 stopes were generated containing 9,393 kt (diluted) with average grades of Au 4.23 g/t, Ag 49.8 g/t, Pb 1.63%, Zn 2.62% and an average calculated NSR of CDN$310/t.

The diluted mill feed derived from measured and indicated resources comes from 794 of the 1579 Main Zone stopes and totals 4,827 kt with an average grade of Au 4.56 g/t, Ag 44.34 g/t, Pb 1.56%, Zn 2.88% and an average calculated NSR of CDN$328/t.

The diluted mill feed derived from inferred resources comes from the remaining 785 of the 1579 Main Zone stopes for a total of 4,566 kt with average grades of Au 3.89 g/t, Ag 55.6 g/t, Pb 1.70%, Zn 2.36% and an average calculated NSR of CDN$291/t.

The undiluted tonnes in the 1579 Main Zone stopes were a total of 6,461 K with average grades of Au 5.93 g/t, Ag 69.48 g/t, Pb 2.28%, Zn 3.68% and an average calculated NSR of CDN$436/t.

The measured and indicated resources in the 794 of the 1579 Main Zone stopes total 3,405 kt with average grades of Au 6.24 g/t, Ag 60.36 g/t, Pb 2.12 %, Zn 3.92% and an average calculated NSR of CDN$449/t.

The inferred resources in the remaining 785 of the 1579 Main Zone stopes total 3,056 kt with average grades of Au 5.58 g/t, Ag 79.63 g/t, Pb 2.45%, Zn 3.41% and an average calculated NSR of CDN$420/t.

The portion of the Main Zone's undiluted mineral resources by class included in the mine plan is shown in Table 16.5.

Resource Class	Tonnes	Avg NSR	Au (g/t)	Ag (g/t)	Pb (%)	Zn (%)
Measured	1,147,241	493	6.71	68.79	2.38	4.38
Indicated	2,257,765	428	6.00	56.07	1.99	3.68
Inferred	3,056,471	420	5.58	79.63	2.45	3.41

Table 16.5 Undiluted Main Zone Mineral Resources included in the Mine Plan

The Main Zone's total diluted mill feed included in the mine plan, is shown in Table 16.6.

Resource Class	Tonnes	Avg NSR	Au (g/t)	Ag (g/t)	Pb (%)	Zn (%)
Measured	1,597,883	365	4.98	51.32	1.77	3.26
Indicated	3,229,183	309	4.35	40.88	1.46	2.68
Inferred	4,565,806	291	3.89	55.58	1.70	2.36

Table 16.6
Diluted Mill Feed from the Main Zone Mineral included in the Mine Plan

16.4.2 Yellowjacket Zone

Mining dilution in the Yellowjacket Zone comprises an average 4.4% allowance for overbreak and a further 34.3% for internal (waste or marginal grade mineralization) included within the stope shapes where operationally necessary, for a total dilution of 38.7% included within the mine schedule.

In the Yellowjacket Zone a total of 78 stopes were generated containing 646,000 t (diluted) with average grades of Au 0.06 g/t, Ag 43.0 g/t, Pb 1.90%, Zn 7.47% and an average calculated NSR of CDN$218/t. All of the diluted mill feed in the 78 Yellowjacket Zone stopes is derived from Indicated Resource.

The undiluted resource portion of these 78 Yellowjacket stopes is all classified as Indicated and has a total of 406 Kt undiluted, with average grades of Au 0.09 g/t, Ag 67.24 g/t, Pb 2.98%, Zn 11.71% and an average calculated NSR of CDN$342/t.

Neither Measured nor Inferred Resources are included in the Yellowjacket mine plan. The portion of the Yellowjacket Zone's undiluted mineral resources included in the mine plan is shown by class in Table 16.7.

Resource Class	Tonnes	Avg NSR	Au (g/t)	Ag (g/t)	Pb (%)	Zn (%)
Measured	0	0	0	0	0	0
Indicated	406,322	342	0.09	67.24	2.98	11.71
Inferred	0	0	0	0	0	0

Table 16.7 Undiluted Yellowjacket Zone Mineral Resources included in the Mine Plan

Table 16.8 shows the total diluted mill feed included in the Yellowjacket Zone mine plan.

Table 16.8 Diluted Mill Feed from the Yellowjacket Zone included in the Mine Plan

Resource Class	Tonnes	Avg NSR	Au (g/t)	Ag (g/t)	Pb (%)	Zn (%)
Measured	0	0	0	0	0	0
Indicated	646,512	218	0.06	43.00	1.90	7.47
Inferred	0	0	0	0	0	0

16.5 EQUIPMENT REQUIREMENTS

The equipment required to meet the production and development schedules are approximated in Table 16.10. The main objectives of mobile fleet will be to consistently achieve the scheduled 2,300 t/d of mill feed which will depend upon keeping and average of at least four stopes producing mill-feed at any given time. A production study to determine cycle times and the exact equipment requirements on a yearly basis are recommended for the next level of study.

Description	Year 0	Year 1-11	Year 12
Production Rate	0 t/d	2,300 t/d	2,300 t/d
Equipment Type	Number	Number	Number
3-Boom Jumbo Drill	3	3*	0
2-Boom Jumbo Drill	5	5	4
Long hole Drill	1	5	5
Bolter	4	4	2
Haul Truck (40-Ton Capacity)	4	4	2
Haul Truck (30-Ton Capacity)	2	6	5
Remote LHD (6 CY Capacity)	1	4	4
LHD (6 CY Capacity)	3	5	3
LHD (3.5 –4 CY Capacity)	3	4	2
Scissor Lift	3	4	3
ANFO Loader	2	4	4
Lube Truck	1	2	2
Crane Truck	1	2	2
Personnel Carrier	1	4	4
Grader	1	2	2
Utility Tractor	1	6	4
Total	36	64	48

Table 16.8 Estimated Mining Equipment Fleet

*After ramp development is completed in year ten, the number of 3-boom drills can be reduced to 1. Note: Some items are currently on site and are sunk costs.

Ventilation raises and ore-passes will be developed during years zero to the end of production year three. The raises will be created using a rented or leased Robbins raise borer (or similar) able to bore 3m diameter raises. The raise boring may be hired out to specialized raise boring contractors.

Power demand for the underground mine is expected to be approximately 32,000 MWh/y. This includes power supply for the shop, yard, underground operations, pump station and compressor. A surface load centre will feed the shop, office, and other miscellaneous surface support facilities.

Underground power supplied at 4,160 V will be distributed through a 1,000 kVA main mine load centre and several 500 kVA load centres. These will run the underground pump station, sumps, fans and other electrically powered equipment.

To achieve the required rates of 16m/day of horizontal and ramp development, operations will require an average of approximately five (5) advances of 3.3 metres (10') per day. Assuming that, for any advancing face, drill and blast will occur on one shift, the second shift will allow the gases to clear, followed by mucking on shift 3, then an average of five jumbo shifts will be needed daily to advance five faces.

To accommodate for mechanical downtime for planned and unplanned maintenance, three larger jumbos are recommended to be assigned to ramp development and five smaller jumbos are assigned to horizontal development, for a total of eight jumbos.

The additional equipment allows for combined mechanical availabilities and equipment utilization in the range of 60-65% (80% mechanical availability x 80% utilization).

The number of LHDs estimated assumes that:

1. 5 development faces will need to be mucked almost daily.
1. 4 stopes in production will need to be mucked daily.

The number of trucks estimated assumes that:

1. Four 30-ton (27 tonne) trucks will be working at any given time producing over a cumulative 1,700 t/d based on an assumed average cycle time of 1 trip per +/- 1 hour and 16+ trips per truck per day.
1. Three 40-ton (36 tonne) trucks will be working at any given time producing over a cumulative 1,600 t/d based on an assumed average cycle time of 1 trip per +/- 1 hour and 16+ trips per truck per day.

The combined capacity will be required to produce mill feed mucked from the stopes and sills as well as waste that will be removed from the mine or moved into empty stopes as backfill.

Current surface facilities (Figure 16.8) include an insulated shop that is suitable for initial mine development activities ( Figure 16.8). As production increases, an underground shop will need to be added to the maintenance facilities.

Figure 16.8 Existing Surface Maintenance Facilities

Figure 16.8 Insulated Shop

Source: Rokmaster video athttps://youtu.be/AWHniPVDrIo

Surface and underground laydown areas will also be needed for storage of consumables and other supplies.

16.6 UNDERGROUND OPERATIONS PERSONNEL

The proposed work schedule for underground operations consists of three eight-hour shifts.

The mine will operate 7 days per week. Management and technical staff will work on the dayshifts only, 5 days per week.

Each piece of production equipment will require three operators on the payroll.

In order to maintain 5 development faces at all times, operations will require three teams of 10 jumbo operators and helpers.

Similarly, in order to maintain 2,300 t/d the operation will require a minimum of 4 stopes in production at any time, consequently 5 long-hole drills will be required with three operators and three helpers per machine for a total of 30 employees.

The number of truck drivers on payroll will also be 30 in total with 10 operating on any given shift.

A total of 30 LHD operators are estimated in order to meet required mucking of stopes and development faces.

A study on optimizing the shift schedules and roster may reveal opportunities to reduce the overall headcount at the next level of study.

Table 16.9 summarizes a high-level estimate of underground personnel that could be required to meet development and production targets in the LOM production schedule.

	Year -1	Year 1-11	Year 12
Description	Number	Number	Number
Shift Supervisors	3	12	12
Jumbo Drillers	30	30	12
Bolters	24	24	12
Long-hole Drillers	3	15	15
Driller's Helpers	3	15	15
Mucker Operators	15	30	24
Truck Drivers	21	30	24
ANFO Truck/blasthole loading	12	24	24
Mine Labor	6	24	24
Mine Superintendent	1	1	1
Clerk	1	2	2
Chief Engineer	1	1	1
Mine Engineer	1	2	2
Surveyor	2	4	2
Survey /Engineering Technicians	2	4	2
Chief Geologist	1	1	1
Geologist	1	4	4
Grade Control Sampler	1	4	4
Maintenance Superintendent	1	1	1
Lead Mechanic	1	2	1
Electrician	1	4	4
Welder	1	4	2
Maintenance Technician	1	8	6
Grader Operator	3	6	6
Drill Mechanic	1	6	6
Mechanic	6	18	12
Total	143	276	219

Table 16.9 Summary of Underground Operations Personnel

16.7 VENTILATION

The ventilation requirements are based upon the maximum estimated amount of equipment that will be in operation in the mine at any given time. The ventilation requirements during years one to nine is estimated at 273m3 /second (576 CFM). This airflow capacity will be provided by two 300 CFM fans that will push air into the mine from near to the main portal. The ventilators will push the air through the air heaters so as work with the warm air's natural tendency to rise up and out of the mine and in order to avoid drawing cold air into the main adit.

Table 16.10 presents the ventilation air requirements over the LOM period.

Table 16.10 Mine Ventilation

	Power	Usage	Year -1	Year 1 -9	Year 10-11	Year 12
Equipment Type	(kW)	Factor	Units	Units	Units	Units
3-Boom Jumbo Drill	175	0.3	3	3	0	0
2-Boom Jumbo Drill	150	0.3	5	5	5	4
Longhole Drill	150	0.5	1	5	5	5
Bolter	55	0.5	4	4	4	2
Haul Truck (20-Ton Capacity)	150	0.75	4	4	4	2
Haul Truck (40-Ton Capacity)	280	0.75	3	6	6	6
Remote LHD (6 CY Capacity)	138	0.75	1	4	4	4
LHD (6 CY Capacity)	138	0.75	2	6	6	4
LHD (3.5 CY Capacity)	115	0.75	2	2	2	2
Scissor Lift	61	0.75	3	4	4	3
Anfo Loader	70	0.5	2	4	4	4
Lube Truck	61	0.5	1	2	2	2
Crane Truck	65	0.5	1	2	2	2
Personnel Carrier	60	0.5	1	4	4	4
Grader	65	0.5	1	2	2	2
Utility Tractor	50	0.5	1	6	6	4
Total			35	63	60	50

	Operating	Factor	Fresh Air	Fresh Air	Fresh Air	Fresh Air
	(kW)	(m3/kW)	(m3/sec)	(m3/sec)	(m3/sec)	(m3/sec)
3-Boom Jumbo Drill	52.5	0.0598	9.4	9.4	0.0	0.0
2-Boom Jumbo Drill	45	0.0598	13.4	13.4	13.4	10.8
Long-hole Drill	75	0.0598	4.5	22.4	22.4	22.4
Bolter	27.5	0.0598	6.6	6.6	6.6	3.3
Haul Truck (20-Ton Capacity)	112.5	0.0598	26.9	26.9	26.9	13.4
Haul Truck (40-Ton Capacity)	210	0.0598	37.6	75.2	75.2	75.2
Remote LHD (6 CY Capacity)	103.5	0.0598	6.2	24.7	24.7	24.7
LHD (6 CY Capacity)	103.5	0.0598	12.4	37.1	37.1	24.7
LHD (3.5 CY Capacity)	86.25	0.0598	10.3	10.3	10.3	10.3
Scissor Lift	45.75	0.0598	8.2	10.9	10.9	8.2
Anfo Loader	35	0.0598	4.2	8.4	8.4	8.4
Lube Truck	30.5	0.0598	1.8	3.6	3.6	3.6
Crane Truck	32.5	0.0598	1.9	3.9	3.9	3.9
Personnel Carrier	30	0.0598	1.8	7.2	7.2	7.2
Grader	32.5	0.0598	1.9	3.9	3.9	3.9
Utility Tractor	25	0.0598	1.5	9.0	9.0	6.0
Total			148.6	272.9	263.5	226.0

16.8 SCHEDULE

Micon has based the Life of Mine Production Schedule upon stopes generated using the P&E Block Model and Datamine's Mineable Shape Optimizer software.

A total of 1579 stopes from the main zone and 78 stopes from Yellowjacket zone were sequenced to generate a schedule that mined levels closest to existing development and the portal first, while also prioritizing stopes with the best gold grades first.

Development headings are prioritized to meet production goals, and mine production is scheduled such that four different areas of the mine can be mined independently.

Higher productivity is realized by effectively dividing the mine into four separate production quadrants, opening multiple development and production headings simultaneously.

The ramping and vertical raises required for production in year one will begin in preproduction year zero (also referred to as Year -1 or Y-1).

The average LOM mining rate is 2,300 t/d.

A summary of the annual mine development and production is shown in the Life-of-Mine Production Schedule (Table 16.11).

Table 16.11 Life-of-Mine Production Schedule

	Unit	Y-1	Y1	Y2	Y3	Y4	Y5	Y6	Y7	Y8	Y9	Y10	Y11	Y12	Total
Ramps (5 m x 5 m)	m	666.7	666.7	400.0	666.7	800.0	400.0	333.3	333.3	333.3	466.7	266.7	0.0	0.0	5,333
Ramps (SG3.7)	kt	61.7	61.7	37.0	61.7	74.0	37.0	30.8	30.8	30.8	43.2	24.7	0.0	0.0	493
Air Raise Development (3m radius)	m	488.4	488.4	488.4	488.4	0.0	0.0	0.0	0.0	0.0	0.0	0.0	0.0	0.0	1,954
Man-ways and Ore-Passes (3 m x 3 m)	m	244.2	244.2	146.5	244.2	293.0	146.5	134.3	134.3	134.3	134.3	97.7	0.0	0.0	1,954
Vertical Development Tonnes (SG3.7)	kt	20.9	20.9	17.7	20.9	9.8	4.9	4.5	4.5	4.5	4.5	3.3	0.0	0.0	116
Dev.- Horizontal (3.5 m x 4 m) - including sills	m	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	5,099.4	66,292
Dev.- Horizontal (3.5 m x 4 m) - without sills	m	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	4,244.0	55,172
Dev.- Horizontal (SG 3.7) - without sills	kt	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	2,858
Waste from Sills	kt	0.0	22.3	30.2	27.1	32.1	43.9	42.6	41.1	52.9	71.2	61.7	62.8	57.8	546
Percent of Mill feed from Sills	%	0	21.3	22.6	22.4	23.5	23.8	23.0	23.0	23.9	24.6	23.2	22.6	21.0	22.9
Mill Feed Tonnes from Sills	kt	0.0	179.1	189.4	188.0	197.1	199.5	192.8	192.7	200.3	206.9	194.9	189.4	169.3	2,299
Mill Feed Tonnes from Stopes	kt	0.0	660.4	650.1	651.5	642.4	640.0	646.7	646.8	639.2	632.6	644.6	650.1	635.6	7,740
Production Mill Feed - Diluted tonnes (Sills+Stopes)	kt	0.0	839.5	839.5	839.5	839.5	839.5	839.5	839.5	839.5	839.5	839.5	839.5	804.9	10,039
NSR	$/t	0	412	417	433	421	385	265	228	241	229	217	216	180	304
Au Grade	g/t	0.00	6.48	6.10	6.10	5.90	5.49	3.46	3.01	2.95	2.67	2.41	2.07	0.85	3.97
Ag Grade	g/t	0.00	47.28	52.22	61.49	62.75	56.55	40.57	34.40	58.88	52.58	39.36	37.89	48.35	49.36
Pb Grade	%	0.00	1.49	1.82	2.09	2.15	1.80	1.45	1.27	1.55	1.58	1.36	1.43	1.76	1.65
Zn Grade	%	0.00	2.43	3.05	3.31	3.15	2.90	2.70	2.33	2.26	2.42	3.02	3.59	4.12	2.94
Vertical Development Rate (350 d/y)	m/d	2.1	2.1	1.8	2.1	0.8	0.4	0.4	0.4	0.4	0.4	0.3	0.0	0.0	0.8
Ramps +Horizontal Dev. Rate (350 d/y) - no sills	m/d	16.8	16.8	16.0	16.8	17.2	16.0	15.8	15.8	15.8	16.2	15.7	14.9	14.9	16.0
Development Rate (350 d/y)	kt/d	0.9	0.9	0.8	0.9	0.9	0.8	0.7	0.7	0.7	0.8	0.7	0.6	0.6	0.8
Mill Feed Production Rate (365 d/y)	kt/d	0.0	2.3	2.3	2.3	2.3	2.3	2.3	2.3	2.3	2.3	2.3	2.3	2.3	2.3
Main Zone - Diluted tonnes (Sills+Stopes)	kt	0.0	839.5	839.5	839.5	839.5	839.5	839.5	839.5	839.5	839.5	677.9	679.0	480.4	9,393
NSR	$/t	0	412	417	433	421	385	265	228	241	229	198	217	177	310
Au Grade	g/t	0.00	6.48	6.10	6.10	5.90	5.49	3.46	3.01	2.95	2.67	2.97	2.54	1.38	4.24
Ag Grade	g/t	0.00	47.28	52.22	61.49	62.75	56.55	40.57	34.40	58.88	52.58	33.42	38.69	56.27	49.80
Pb Grade	%	0.00	1.49	1.82	2.09	2.15	1.80	1.45	1.27	1.55	1.58	1.02	1.41	1.83	1.63
Zn Grade	%	0.00	2.43	3.05	3.31	3.15	2.90	2.70	2.33	2.26	2.42	1.40	2.67	2.67	2.62
Yellowjacket - Diluted tonnes (Sills+Stopes)	kt	0.0	0.0	0.0	0.0	0.0	0.0	0.0	0.0	0.0	0.0	161.6	160.5	324.4	647
NSR	$/t	0	0	0	0	0	0	0	0	0	0	297	208	184	218
Au Grade	g/t	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.05	0.06	0.06	0.06
Ag Grade	g/t	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	64.25	34.50	36.63	43.00
Pb Grade	%	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	2.77	1.53	1.66	1.90
Zn Grade	%	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	0.00	9.85	7.49	6.27	7.47
Waste Summary	mass/length	Y-1	Y1	Y2	Y3	Y4	Y5	Y6	Y7	Y8	Y9	Y10	Y11	Y12	Total
Source	t/m	kt	kt	kt	kt	kt	kt	kt	kt	kt	kt	kt	kt	kt	kt
Ramps (5m x5m x SG3.7)	92.5	61.7	61.7	37.0	61.7	74.0	37.0	30.8	30.8	30.8	43.2	24.7	0.0	0.0	493
Vertical Development Tonnes (SG 3.7)	26 to 33	20.9	20.9	17.7	20.9	9.8	4.9	4.5	4.5	4.5	4.5	3.3	0.0	0.0	116
Dev.- Horizontal (3.5m x4m x SG 3.7)	52	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	219.8	2,858
Waste from Sills (external to stopes < 5m in width)	92.5	0.0	22.3	30.2	27.1	32.1	43.9	42.6	41.1	52.9	71.2	61.7	62.8	57.8	546
Total Waste	kt	302	325	305	330	336	306	298	296	308	339	310	283	278	4,013

17.0 RECOVERY METHODS

The preliminary process plant flowsheets and design criteria were developed using the results from the metallurgical testwork programs that are discussed in Section 13.0 of this Technical Report.

The base case flowsheet for the Main Zone mineralization comprises primary and secondary crushing, pre-concentration using heavy media separation (HMS), grinding, gravity separation, sequential flotation of lead, zinc and gold rich sulphide concentrates, concentrate dewatering, and tailings disposal. The base case also includes the pressure oxidation of the gold rich sulphide concentrate and subsequent gold recovery using carbon-in-leach (CIL) technology, while the alternative case assumes that the gold rich sulphide concentrate is sold to an international smelter/refinery.

Simple process block flow diagrams illustrating the selected Main Zone process flow and gold recovery are presented in Figure 17.1 and Figure 17.2, respectively.

The flowsheet for the processing of Yellowjacket Zone mineralization is similar to the Main Zone base case except that the HMS circuit, gold rich sulphide flotation and POX circuits are bypassed. The gold recovery circuit is not required when treating Yellowjacket Zone material.

17.1 PROCESS DESIGN CRITERIA

The process design is based on a nominal plant feed rate of 2,300 t/d. The process design basis is provided in Table 17.1 and the process design criteria are summarized in Table 17.2 and Table 17.3. The design criteria are based upon the metallurgical testwork which was discussed in Section 13.0 of this Technical Report, standard industry design factors, and Micon's inhouse experience.

17.1.1 Design Basis

Table 17.1 summarizes the process design basis. The operation is designed to treat 2,300 dry t/d of mineralization from the underground mine on the basis of a 24 hour per day, 7 day per week operation.

The utilization factors used for the calculation of the nominal hourly flow rates are 65% for the primary and secondary crushing circuit and 92% for the remainder of the mineral processing facilities and the gold recovery plant. These factors will be reviewed with equipment suppliers during the next phase of project development.

Figure 17.1 Main Zone Mineral Processing Block Flow Diagram

Figure 17.2 Main Zone Gold Recovery Block Flow Diagram

Table 17.1 Process Design Basis

Parameter	Units	Main	Yellowjacket	Source
Operating time	days/year	365	365	Micon
Operating time	hours/day	24	24	Micon
Crushing operating criteria	days/week	7	7	Micon
Crushing utilization	%	65	65	Micon
Plant operating criteria	days/week	7	7	Micon
Grinding & flotation plant utilization	%	92	92	Micon
Gold plant utilization		92	na
Throughput
Nominal annual throughput	kt	365	365	Micon
Design daily throughput	t	2,300	2,300	Micon
Run-of-Mine Ore Characteristics (plant design only)
Maximum rock size	mm	500	500	Micon
Ore specific gravity		3.29	2.95	Micon
Ore moisture	wt %	5.0	5.0	Micon
Design feed grade –Pb	%	1.80	1.90	Average June 2020 Mine
				Plan 1st 3 years
Design feed grade –Zn	%	2.93	7.47	Average June 2020 Mine
				Plan 1st 3 years

Parameter	Units	Main	Yellowjacket	Source
Average feed grade –Au	g/t	6.23	0.06	Average June 2020 Mine
				Plan 1st 3 years
Average feed grade –Ag	g/t	53.66	43.00	Average June 2020 Mine
				Plan 1st 3 years
Metallurgical Efficiency (plant design only)
HMS –Loss to floats	wt%	28.0	na	Testwork
HMS –Loss to floats	Pb%	1.0	na	Testwork
HMS –Loss to floats	Zn%	2.3	na	Testwork
HMS –Loss to floats	Au%	0.5	na	Testwork
HMS –Loss to floats	Ag%	1.7	na	Testwork
Recovery to gravity concentrate	wt%	4.9	na	Testwork
Recovery to gravity concentrate	Pb%	11.7	na	Testwork
Recovery to gravity concentrate	Zn%	5.3	na	Testwork
Recovery to gravity concentrate	Au%	30.2	na	Testwork
Recovery to gravity concentrate	Ag%	10.1	na	Testwork
Gravity concentrate grade	Au g/t	36.4	na	Testwork
Gravity concentrate grade	Ag g/t	141	na	Testwork
Recovery to lead concentrate	Pb%	82.8	87.5	Testwork
Recovery to lead concentrate	Zn%	8.6	5.4	Testwork
Recovery to lead concentrate	Au%	11.6	58.3	Testwork
Recovery to lead concentrate	Ag%	65.5	81.9	Testwork
Lead concentrate grade	Pb%	54.3	43.3	Testwork
Recovery to zinc concentrate	Pb%	8.8	6.7	Testwork
Recovery to zinc concentrate	Zn%	81.0	93.1	Testwork
Recovery to zinc concentrate	Au%	1.7	27.5	Testwork
Recovery to zinc concentrate	Ag%	9.9	12.3	Testwork
Zinc concentrate grade	Zn%	55.0	61.9	Testwork
Recovery to gold concentrate	Au%	85.7	na	Testwork
Recovery to gold concentrate	Ag%	23.6	na	Testwork
Gold flotation concentrate grade	Au g/t	17.0	na	Testwork
Gold flotation concentrate grade	Ag g/t	70.3	na	Testwork
Lead concentrate production –nominal	dry t/d	55	88	Calculation
Lead concentrate production -design	dry t/d	68	111	Calculation
Zinc concentrate production –nominal	dry t/d	91	258	Calculation
Zinc concentrate production -design	dry t/d	114	323	Calculation
Gold concentrate production –nominal	dry t/d	590	na	Calculation
Gold concentrate production -design	dry t/d	708	na	Calculation

Notes: Gravity recoveries based on plant feed.

Flotation recoveries based on flotation feed.

17.1.2 Process Design Criteria

Table 17.2 and Table 17.3 provide an overview of the process design criteria for mineral processing and gold recovery, respectively.

A process flow diagram (PFD) showing the major unit operations was prepared by Canenco and is presented as Figure 17.3.

Description	Units	MainZone	YellowjacketZone
General Feed Characteristics
Annual Processing Rate	t/y	840,000	840,000
Operating Days Per Year	d/y	365	365
Daily Processing Rate	t/d	2,300	2,300
Crusher work index (Cwi)	kWh/t	9.7 -12.7
Bond ball mill work index (metric)	kWh/t	9.5	9.8
Specific gravity		3.29	2.95
Crushing
Crusher Rate -Nominal	t/h	147	147
Primary Crusher Type		Jaw	Jaw
Primary Crusher Size	mm	840 x 1045	840 x 1045
Primary Crusher Closed-Side Setting	mm	75	75
Secondary Crusher Type		Standard	Standard
Secondary Crusher Size	mm	830	830
Secondary Crusher Closed-Side Setting	mm	25	25
Secondary Screen Passing Size	mm	25	25
Crushed Ore Product P80	mm	18.3	18.3
Heavy Media Separation
Heavy Media Circuit Availability	%	90	-
Feed Rate to Wet Screen -Design	t/h	106.5	-
Wet Screen Mesh Size	mm	2.0	-
Percent Wet Screen Fines	%	8.0	-
Wet Screen Fines to Grinding	t/h	7.7	-
Feed Rate to Heavy Media Vessel	t/h	98.8	-
HMS Float Sink (Product)	% wt. of	72	-
HMS Float Reject	% wt. of	28	-
HMS Float Reject	t/h	29.6	-
HMS Media Type		Ferrosilicon	-
HMS Media SG		2.85	-
HMS Sinks SG		3.90	-
HMS Floats SG		2.77	-
Grinding
Total Feed Tonnage to Grinding -Nominal	t/d	1,660	2,300
Feed Rate to Grinding -Nominal	t/h	75.2	104.2
Grinding Circuit Feed F80	mm	17.500	17.500
Rod Mill Discharge P80	micron	800	-
Rod Mill Unit Power Consumption	kWh/t	5.36	-
Rod Mill Installed Power	kW	400	-
Ball Mill Cyclone Overflow P80	micron	34	74
Ball Mill Unit Power Consumption	kWh/t	13.57	9.67
Ball Mill Installed Power	kW	1,200	1,200
Gravity Separation
Feed Rate to Gravity Circuit –Design	dry t/d	1,660	-
Final Concentrate Recovered	dry t/d	116	-
Lead Flotation and Regrinding
Feed Rate to Lead Flotation –Design	dry t/h	69.9	104

Table 17.2 Process Design Criteria – Mineral Processing

Description	Units	MainZone	YellowjacketZone
Pb Rougher Retention Time	min	38	34
Pb Rougher Cell Volume	m3	130	119
Pb Scavenger Retention Time	min	24	-
Pb Scavenger Cell Volume	m3	74	-
Pb Regrind Circuit Feed Tonnage -Nominal	dry t/h	8.2	12.0
Pb Regrind Mill Unit Power Consumption	kWh/t	12.5	13.3
Pb Concentrate Regrind COF P80	micron	15	23
Pb Concentrate Regrind Installed Power	kW	224	224
Pb 1st Cleaner Retention Time	min	43	38
Pb 1st Cleaner Scavenger Retention Time	min	20	16
Pb 2nd Cleaner Retention Time	min	32	14
Pb 3rd Cleaner Retention Time	min	20	10
Pb 4th Cleaner Retention Time	min	18	-
Lead Concentrate Mass Pull of Float Feed	%	3.6	3.8
Lead Concentrate Production –Annual	t/y	20,075	32,300
Lead Concentrate Production –Nominal	t/d	55	88
Lead Concentrate Grade	%Pb	54.3	43.3
Lead Concentrate Recovery	%Pb	82.8	87.5
Zinc Flotation
Feed Rate to Zinc Flotation -Nominal	t/h	67.4	100
Zn Rougher Retention Time	min	24	30
Zn Rougher Cell Volume	m3	96	125
Zn Scavenger Retention Time	min	14	-
Zn Scavenger Cell Volume	m3	45	-
Zn Regrind Feed Tonnage -Nominal	t/h	20.6	19.6
Zn Regrind Mill Unit Power Consumption	kWh/t	12.5	7.6
Zn Concentrate Regrind COF P80	micron	15	35
Zn Concentrate Regrind Installed	kW	337	337
Zn 1st Cleaner Retention Time	min	30	25
Zn Cleaner Scavenger Retention Time	min	13	12
Zn 2nd Cleaner Retention Time	min	21	8
Zn 3rd Cleaner Retention Time	min	15	5
Zn 4th Cleaner Retention Time	min	12	-
Zn Concentrate Mass Pull of Float Feed	%	5.9	11.2
Zn Concentrate Production –Annual	t/y	33,200	94,300
Zn Concentrate Production –Nominal	t/d	91	258
Zn Concentrate Grade	%Zn	55.0	61.9
Zn Concentrate Recovery	%Zn	81.0	93.1
Pyrite-Gold Flotation
Feed rate to Gold Flotation-Nominal	t/h	63.3	-
Au Rougher Retention Time	min	60	-
Au Rougher Cell Volume	m3	176	-
Au Scavenger Retention Time	min	38	-
Au Scavenger Cell Volume	m3	67	-
Au 1st Cleaner Retention Time	min	20	-
Au Concentrate Mass Pull of Float Feed	%	31	-
Au Concentrate Production –Annual	t/y	173,000	-
Au Concentrate Production –Nominal	t/d	474	-
Au Concentrate Grade -Nominal	g/t Au	17.0	-

Description	Units	MainZone	YellowjacketZone
Au Concentrate Recovery	%Au	85.7	-
Flotation Tailings -Annual	t/y	342,000	713,000
Flotation Tailings Production -Nominal	t/d	937	1,953
Concentrate Dewatering
Pb Concentrate Solids Feed Rate	t/h	2.48	2.38
Pb Concentrate Thickener Unit Area	t/m2/h	0.6	0.6
Pb Concentrate Thickener Diameter	m	3.5	3.5
Pb Concentrate Thickener u/f –solids density	wt% solids	60	60
Pb Concentrate Unit Filtering Rate	kg/m2/h	250	250
Pb Concentrate Pressure Filter Area	m2	25	31
Pb Concentrate Filter Cake –solids density	wt% solids	8	8
Zn Concentrate Solids Feed Rate	t/h	4.13	11.7
Zn Concentrate Thickener Unit Area	t/d/m2	0.6	0.8
Zn Concentrate Thickener Diameter	m	5.0	5.0
Zn Concentrate Thickener u/f –solids density	wt% solids	60	60
Zn Concentrate Unit Filtering Rate	kg/m2/h	250	250
Zn Concentrate Pressure Filter Area	m2	28	150
Zn Concentrate Filter Cake –solids density	wt% solids	8	8
Au Concentrate Solids Feed Rate	t/h	25.6	-
Au Concentrate Thickener Unit Area	t/d/m2	0.8	-
Au Concentrate Thickener Diameter	m	7.0	-
Au Concentrate Thickener u/f –solids density	wt% solids	65	-
Tailings Dewatering
Tailings Solids Feed Rate	t/h	60.1	88.5
Tailings Thickener Unit Area	t/m2/h	0.42	0.83
Tailings Thickener Diameter	m	15	13
Tailings Thickener u/f –solids density	wt% solids	55	55
Tailings Unit Filtering Rate	kg/m2/h	185	185
Tailings Filter Filter Area	m2	400	480
Tailings Filter Cake –solids density	wt% solids	15	15

Table 17.3 Process Design Criteria – Gold Recovery Plant

Description	Units	Main Zone
Pressure Oxidation Feed Characteristics
Annual Processing Rate	t/y	215,300
Daily Processing Rate	t/d	590
Average Feed Grades	%Pb	1.2
	%Zn	1.5
	g/t Au	22
	g/t Ag	65
	%As	17
	%S	36
	%Fe	34
Specific Gravity	t/m3	3.29
Circuit Availability	%	90
Oxygen Pressure, over steam	kPa	700
Pressure Oxidation Feed Pulp Density	% solids	50

Description	Units	Main Zone
Design retention Time	h	1.0
Autoclave Temperature	°C	200
Total Operating Pressure	kPa	2827
Design Operating Pressure	kPa	3200
Design Temperature	°C	235
Slurry Feed % solids	%	20
Oxygen Demand, t O2/t feed	%	50
Oxygen Demand tonne/day –Design	t/d	283
Feed Solids Leached (Mass)	%	15
Autoclave Discharge
Residue Free Acid (H2SO4) Content	g/L	50
Autoclave Discharge Conditioning Time	h	5
Autoclave Discharge CCD Circuit -Stages	-	5
Autoclave Discharge CCD Circuit Wash Ratio	-	2
Lime Boil	-	Not Included
Solution Neutralization
Flow rate	m3/h	109
Retention time	h	3
Limestone addition	kg/t solids	400
First Stage pH		1.2
Lime addition	kg/t solids	30
Second Stage pH		3.4
Thickener circuit feed -solids	t/d	542
Thickener u/f density	wt%	55
Thickener sizing criteria	t/m2/h	1.35
Thickener diameter	m	7.0
Gold Leaching Circuit
Feed Rate to CIP Circuit-Nominal	t/d	501
Feed Rate to CIP Circuit -Nominal	t/h	22.7
Feed Rate to CIP Circuit -Nominal	m3/h	49.1
Leach Slurry Density	% solids	40
Leach Residence Time	h	24
Number of Leach Tanks	#	6
Carbon Density	g/L	7.0
Loaded Carbon Gold Grade	g Au/t	3,500
Loaded Carbon Production	t/d	2.4
Gold Extraction, Based on Autoclave Feed	%	95
Silver Extraction, Based on Autoclave Feed	%	35
Gold Produced -Nominal	oz/y	142,000
Silver Produced -Nominal	oz/y	157,000
Doré Produced -Nominal	oz/d	909
Effluent Treatment
Feed Rate to Cyanide Destruction -Nominal	m3/h	49.1
Cyanide Destruction Residence Time	h	2
No. Cyanide Destruction Reactor Tanks		2

Figure 17.3 Proposed Process Flow Diagram

17.2 PROCESS DESCRIPTION

17.2.1 Crushing

The mineralized material will be delivered by mine trucks to the plant and fed through a static grizzly with 500 mm openings to the ROM feed hopper. The material will be extracted from the feed hopper by a grizzly feeder and fed to a 1045 x 840 jaw crusher. The crushed ore with an 80% passing size (P80) of 80 mm will be transported by conveyor to a coarse rock storage bin from which it will be fed to a triple deck vibrating screen. The screen undersize (+ 25 mm) will be fed to the heavy medium separation circuit while the oversize material will feed the secondary cone crusher. Secondary crusher product (P80 28 mm) will be recycled to the classifying screen.

17.2.2 Heavy Media Separation

The heavy medium separation (HMS) circuit will be utilized to upgrade the mineralization from the Main Zone by the selective removal of crushed non-sulphide, relatively light, gangue material. Testwork has suggested that typically about 28% of the feed can be removed as waste with minimal loss of valuable metals. For the PEA a modular 100 t/h HMS cyclone plant with a SG cut-point of 2.85 using ferrosilicon is assumed. Testwork has suggested that good upgrading with minimal losses can be achieved with a crush size of 50 mm and therefore a static bath vessel, such as a drum, could be utilized for this application. However, separation efficiency at smaller particle sizes may be adversely affected as compared to an HMS cyclone.

The HMS plant is fed with crushed material from the storage bin by variable-speed belt feeders discharging onto conveyors, where the material is measured by weightometer and tramp metal is removed by an electromagnet. The crushed mineralization is fed onto a fine removal screen where – 2 mm material is recovered and pumped to the grinding circuit. The coarse (+2 mm) material will feed the HMS feed box where it will be mixed with heavy medium (ferrosilicon slurry of controlled density) and pumped to the HMS cyclone.

The two products from the HMS cyclone, separation sinks (cyclone underflow) and floats (cyclone overflow), will be directed to their respective drain and rinse screens for removal and recovery of ferrosilicon. The ferrosilicon slurry is collected from both screens and directed to the dilute medium tank from which it is pumped to a magnetic separator to recover ferrosilicon.

A densifying circuit consisting of a circulation pump and pipe densifier that will bleed of water from the ferrosilicon circuit. To balance this bleed and to maintain medium density, a density controller will regulate water addition to the circulating medium circuit.

Washed float material will be stockpiled outside the plant area. This material will be reclaimed by a loader and dump trucks and used for construction, reclamation, or backfill purposes.

Sink material, containing relatively heavy sulphide mineralization, will be conveyed to the grinding circuit feed storage bins.

The HMS circuit would be by-passed when processing Yellowjacket material.

17.2.3 Grinding

The HMS product will be fed from the feed bins by two belt feeders at a controlled rate and fed via a conveyor to a 3.05 m diameter by 4.57 m long rod mill. The rod mill discharge, with a P80 of around 800 microns (µm), will discharge into a cyclone feed pump box together with the ball mill discharge. The combined slurry will be pumped to cyclones, the cyclone underflow will feed the ball mill grinding circuit while the cyclone overflow will be directed to the lead flotation circuit.

The 3.66 m diameter by 5.49 m long ball mill will have a circulating load of 300%.

The target cyclone overflow P80 size for Main Zone mineralization is 34 µm and 74 µm for Yellowjacket material. No major changes to the main equipment will be required when the plant feed changes from Main to Yellowjacket material, which is late in the Project life, although adjustments will be made to the feed systems and classifying cyclones.

17.2.4 Gravity Separation Circuit

A portion of the ball mill cyclone underflow will be directed to a gravity feed vibrating screen that will remove coarse material from the process stream to protect the centrifugal gravity separators. The gravity concentrate will be pumped to the gold concentrate storage tank while the gravity tailings will be recycled to the grinding circuit.

17.2.5 Lead Flotation

Ball mill cyclone overflow will be is fed to the lead flotation circuit conditioning tanks where reagents will be added, prior to feeding the lead rougher flotation cells.

The lead flotation circuit will comprise roughing, scavenging, regrinding of combined rougher and scavenger concentrate, four stages of cleaning and cleaner scavenging. The target product size from the regrinding circuit is P80 of 15 µm for Main Zone and 23 µm for Yellowjacket.

Flotation collectors (Aero241, 3418A, PEX) and zinc depressant cyanide/zinc sulphate will be added to the roughers, cleaners and regrind circuit. MIBC (methyl isobutyl carbinol) will be used as a frother and lime as a pH modifier.

The final lead concentrate from the fourth stage of cleaning will be dewatered using a thickener and pressure filter then stored prior to being loaded onto trucks and transported off-site to market.

17.2.6 Zinc Flotation

The tailings from the lead circuit will feed the zinc circuit conditioner which will feed the zinc flotation roughers.

Similar to the lead circuit, the zinc flotation circuit will comprise roughing, scavenging, regrinding of combined rougher and scavenger concentrate, four stages of cleaning and cleaner scavenging. The target product size from the regrinding circuit is P80 of 15 µm for Main Zone and 35 µm for Yellowjacket.

Flotation collector (SIPX), activator (copper sulphate) and pH modifier (lime) will be added to the roughers, cleaners and regrind circuit. MIBC (methyl isobutyl carbinol) will be used as a frother.

The final concentrate from the fourth cleaner will be pumped to the dewatering circuit.

The cleaner-scavenger tails are combined with the zinc rougher tails and pumped to the tailings pond.

17.2.7 Bulk Sulphide-Gold Flotation

The scavenger tailings from the zinc circuit will feed the bulk sulphide flotation circuit that will recover the remaining gold into a sulphide concentrate containing mainly pyrite and arsenopyrite. This circuit will only operate when Main Zone mineralization feeds the plant, it will not be used for Yellowjacket feed.

The gold/bulk sulphide flotation circuit will comprise roughing, scavenging and one stage of cleaning.

Flotation collector (PAX), activator (copper sulphate) and pH modifier (sulphuric acid) will be added to the roughers, cleaners and regrind circuit. MIBC (methyl isobutyl carbinol) will be used as a frother.

The gold concentrates from the cleaner flotation circuit and the gravity circuit will be dewatered and either shipped to market or (in the base case) fed to the gold recovery pressure oxidation circuit.

17.2.8 Dewatering

Lead concentrate will be thickened to about 60% solids by weight in a 3.5 m diameter thickener. Thickener overflow will be recycled to the grinding process water circuit while underflow will be filtered using pressure filters and stored in the concentrate storage area.

Zinc concentrate will be thickened to about 60% solids by weight using a 3.5 m diameter thickener. A de-aeration sump will be located immediately before the zinc thickener to promote

breakdown of the froth and improve the effectiveness of flocculation and subsequent settling within the thickener. Thickener overflow will be recycled to the zinc process water circuit while underflow will be filtered using pressure filters and stored in the concentrate storage area.

Gold concentrate will be thickened to between 50% and 60% solids by weight using a 8.0 m diameter thickener. Thickener overflow will be recycled to the bulk sulphide process water circuit while underflow will be either be filtered using pressure filters and stored in the concentrate storage area or (in the base case) pumped to the gold recovery pressure oxidation circuit.

Tailings will be thickened to 55% solids by weight using a 13.0 m diameter thickener then filtered to approximately 15% moisture and will either diverted to the paste backfill plant or transported to the dry-stack tailings management facility (WMF).

17.2.9 Pressure Oxidation

Pressure oxidation (POX) of the bulk sulphide concentrate will be carried out in an autoclave, using oxygen produced in a dedicated oxygen plant on site. The resulting autoclave discharge slurry will be conditioned, then dewatered with the solids feeding cyanidation and the solution neutralization.

Neutralization will be completed in two stages; the initial stage will use ground limestone and the second will use lime. The precipitated solids from neutralization will be dewatered and stored in a managed facility on site.

17.2.10 Gold Leaching and Recovery

A six-stage CIL circuit will be used to provide 24 hours of leaching time. Lime will be added to maintain a pH of around 10 and sodium cyanide solution will be added to a number of tanks to ensure a concentration of 0.5 g/L NaCN. The activated carbon will be transferred counter current to the process flow and 2.5 t/d of loaded carbon will be transferred to the elution circuit for acid washing, stripping and regeneration. Pregnant solution will feed the electrowinning circuit and electrowinning gold sludge will be recovered and smelted into doré bars containing approximately 43% gold and 47% silver. Average doré production will be about 28 kg per day.

The tailings from the final CIL tank, will gravitate to the carbon safety screen to recover fine carbon particles from the tailings and minimize gold losses. The carbon safety screen undersize will feed the cyanide destruction circuit then be pumped to the tailings dewatering section of the plant.

17.2.11 Reagents

Storage facilities and package plants will be provided to mix and supply the following reagents required for the process:

Potassium Ethyl Xanthate (PAX).
Sodium Isopropyl Xanthate (SIPX).
Potassium Amyl Xanthate (PAX).
AEROFLOAT 241 (Promoter).
AEROPHINE 3418A (Promoter).
AERO 7261A (Pyrite Depressant).
MIBC.
Copper sulphate.
Zinc sulphate
Flocculant.
Sodium cyanide.
Sulphuric acid.
Lime.

Crushed limestone will be stored separately and ground before being added to the neutralization circuit.

17.2.12 Services

Dedicated compressors will be used to generate compressed air for filter, instrument and maintenance purposes. Blowers will produce low pressure air for the flotation process.

In order to minimize the detrimental effect of recycled reagents in the process water there will be multiple process water systems. Circuits that will have separate systems will include HMS, grinding and lead flotation, zinc flotation, bulk sulphide/gold flotation and the gold recovery plant.

To the greatest extent possible the process plant will re-use process water recovered from the WMF to meet process plant demands. Fresh water will only be used where water quality with low dissolved solids is required and as make-up in the process water circuits if mine dewatering water is not of sufficient quality.

17.3 PROCESSING PLANT LAYOUT

Mineralized material will be hauled in trucks from the underground mine to the crusher feed bin sited west of the 832 adit portal. Material will then move southward through the grinding, flotation and dewatering plant, with tailings gravity fed from the west side of the plant into the proposed waste management facility. Pressure oxidation and gold room are located on the south end of the plant area. See Figure 17.4 (over).

Figure 17.4 Proposed Process Plant Layout

Source: Canenco

18.0 PROJECT INFRASTRUCTURE

18.1 ACCESS

Vehicle access to the area is via Provincial Highway 23, approximately 32 km north of the town of Revelstoke, where Highway 23 intercepts the Carnes Creek Forest Service Road. The Property is then reached by travelling eastward 13 km along the Carnes Creek Forest Service Road before reaching the Revel Ridge mine camp.

18.2 POWER SUPPLY

Electric power is currently produced by on-site diesel generators and a VOIP/cellphone and internet system is in place. In the PEA, a provision is made for a transmission line that will tie into the Revelstoke Dam utilities, on-site substation and emergency generators.

The total power demand of the mine, concentrator and gold recovery plant is estimated to be approximately 14 MW and requires a substation capacity of approximately 18 MW.

18.3 WATER SYSTEMS

18.3.1 Process Water

Process water will be reclaimed from the water management pond and pumped back to the plant. As described in Section 17.0, there will be multiple process water systems within the plant to minimise inter circuit reagent contamination.

Mine water will be recycled and used underground for drilling, dust suppression, and maintenance needs. All mine water will report to a main sump underground.

18.3.2 Fresh Water

Run-off will be directed by cut-off ditches to a Fresh Water storage pond located to the west of the process plant. The pond will be maintained at a certain level to provide fire water. Should run-off be insufficient and the pond level decrease, pumps will supply water from nearby continuous streams or wells.

18.3.3 Potable Water

An existing modular potable water packaged plant will be used to provide potable water for the operation.

18.4 FUEL STORAGE

A diesel storage tank will be required at the mine yard and, as the mine continues to develop, underground diesel storage tanks will be located in the underground shop and other locations in the mine as needed.

A storage tank and distribution pumps are required for Heavy Fuel Oil (HFO) that will be used in the carbon regeneration kilns.

All fuel storage tanks will be located in non permeable containment berms having 110% of the capacity of the storage tank.

18.5 BUILDINGS

Existing buildings will be used for the new mine. The cafeteria, administration, warehouse/maintenance shop were confirmed to be in suitable condition. A new change facility is required, and surface facilities will be expanded as the development of the Project ramps up.

For construction, the local operators will provide accommodation for their workers.

18.6 SEWAGE TREATMENT

The existing sewage treatment is currently operating. It has been assumed that these facilities will need to be upgraded to accommodate operations ramp up.

18.7 FIRE PROTECTION

A fire protection system will need to be installed. Firewater pumps are provided in this study.

18.8 TAILINGS DISPOSAL

Thickened tailings will either be pumped underground to be used as cemented paste backfill or filtered and then transported to the WMF for compaction and storage. Water from the WMF will report to the water management pond and be pumped back to the process plant.

18.9 VENTILATION

Ventilation of the mine will be facilitated by two 250,000 CFM (127.5 m3 /sec) fans to push clean air through the heating unit and into the mine.

The ultimate sizing of the primary and secondary fans will be based upon the maximum number of diesel equipment and persons that will be working in the mine at once.

The ventilation arrangement will be designed so as to avoid drawing cold air into the main portals, and to assist the naturally buoyant warm air to rise by convection through the ventilation raises that will daylight on the mountainside above the portal.

18.10 WASTE DUMP

During the mine's 12-year life a total of 4.01 Mt of waste rock will be generated. Some will be used as unconsolidated rock backfill, some will be combined with cemented back fill made primarily from coarse tailings, while the balance will primarily report to the waste dump outside of the mine.

In total:

1. 493 kt of waste will be generated from ramps development.
1. 116 kt of waste will be generated from vertical (sub-vertical) raises.
1. 2,858 kt of waste will be produced from horizontal developments.
1. 546 kt will be the portion of waste generated in sills.

If 100% of the waste rock were to be placed in the waste dump the required waste dump volume needed would be 1.47 Mm3 .

Using a bank density of 3.0 SG for the in-situ waste and a swell factor of 35% gives the waste rock a loose density of approximately 2.22 SG. Consequently, the maximum waste dump size that will need to be designed would contain a volume of 1.47 Mm3 .

A study to determine the quantity of waste rock that could be used as back fill will help to determine the required volume in the waste dump design and the required throughput capacity of the cemented backfill plant.

18.11 CEMENTED BACKFILL PLANT

A cemented backfill plant will have to be sized and designed in order to generate enough backfill from tailings and added sand in enough quantity to backfill behind the 2,300 t/d (622 m3 of empty stopes per day) of production.

The backfill plant will be located to maximize the assistance of gravity and the minimization of pumping for the distribution of the cemented backfill.

18.12 EXPLOSIVES STORAGE

A bulk powder magazine will be located on the site at a location that will meet all regulatory requirements. There will be one bunker for bulk explosives and a second bunker for high explosives (such as blasting caps and boosters) and consumables such as firing line, nonel cords and delays. An underground powder magazine will also be created for storage of explosives needed in the immediate future.

18.13 MINE DEWATERING AND SEDIMENTATION PONDS

Mine dewatering above the 830 m Level will be facilitated by gravity while levels below the 830 m Level will be dewatered using a system of sumps and submersible pumps.

Allowances are made for dewatering sumps to be located on every level below the 830 m elevation and every fourth level above the 830 m elevation.

Submersible trash pumps situated within each of the sumps will be activated by float switches so as to run only when needed. The size and specifications of the pumps will be determined based on ground water inflow prior to and during operations.

Excess mine water that is not kept within a storage tank nor re-used for mining will report to a sedimentation (settling) pond outside of the mine via an HDPE pipe that will exit at or near to the portal at the 832 m level.

The sedimentation pond will be designed to allow for the required retention time so that suspended solids are given adequate time to settle out, so that any effluent reporting to nature will meet applicable environmental regulations. A dosing station may be needed at the sedimentation pond to permit treatment of the water before it exits the pond or is pumped to the process plant.

18.14 MAINTENANCE SHOPS, LAYDOWN AND STAGING AREA

A flat open area 20m x 20 m will be created to allow for receiving, assembling, and commissioning equipment for underground operations and for the processing plant.

The area should be compacted, surfaced and graded to shed rainwater quickly, prevent pooling and to permit easy snow removal with a grader, snowplow or other equipment without damaging its surface.

An allowance of 20 m of development per five levels is included in the mine plan. These underground openings will be used for maintenance workshops of various types, welding and tooling shops, and as storage areas for parts, ground support equipment, and other consumables.

18.15 WASTE MANAGEMENT FACILITY

18.15.1 Basis of Design

Knight Piésold Ltd. (KP) was retained by Rokmaster to develop a conceptual level waste management plan and capital cost estimate to store 4.84 M tonnes of filtered tailings, 2.81 M tonnes of coarse rejects and 3.36 M tonnes of waste rock in a Waste Management Facility (WMF).The key design criteria are provided in Table 18.1 (over).

Item	Criteria
CDA Hazard Classification	Very High
Tailings Placed Dry Density	1.6 t/m3
Coarse Rejects Placed Dry Density	1.8 t/m3
Waste Rock Placed Dry Density	1.88 t/m3
Tailings Production (Years 1 -2)	375,000 m3
Tailings Production (Years 3 –12)	2.65 M m3
Paste Backfill (Tailings) Percentage	36%
Coarse Rejects Production (Years 1 –2)	262,000 m3
Coarse Rejects Production (Years 3 –12)	1.30 M m3
Waste Rock Production (Year 0)	130,000 m3
Waste Rock Production (Years 1 -12)	1.66 M m3
PAG Waste Rock Percentage	5%
Waste Rock Backfill Percentage	5%
Total Tailings Volume on Surface	1.94 M m3
Total Coarse Reject Volume on Surface	1.56 M m3
Total Waste Rock Volume on Surface	1.70 M m3
Water Management Criteria	Collect seepage from the WMF underdrain and
	convey it to a water treatment facility, as required•
	Provide temporary storage of contact water in the
	Water Management Pond ("WMP") based on the
	estimated runoff from the 1 in 10-year, 24-hour
	storm event
•	Convey contact and non-contact water to the
	environment based on the estimated runoff from the
	1 in 200-year, 24-hour storm event

Table 18.1 Waste Management Facility Key Design Criteria

18.15.2 Waste Management Facility Layout

The WMF is designed for incremental development, to contain both the tailings and the waste rock generated through life of mine. The final arrangement of the WMF, along with a typical section, are provided in Figure 18.1.

Figure 18.1 A Plan View and Cross-Section of the Conceptual WMF

Source: Knight Piésold, 2020.

The location of the facility was previously selected by Rokmaster for its proximity to the plant site and gentle topography. The conceptual level layout of the WMF was developed based on the presence of two creeks that pass through the property to the north and west of the plant site, and the more rugged topography to the east.

18.15.3 Waste Management Facility Construction

The Stage 1 WMF was developed to provide two years of waste storage. The Stage 1 WMF will be approximately 15 m in height with the Perimeter Zone (shell), coarse rejects and tailings constructed to crest El. 795 m. The overall outer slope angle of the WMF will be 3.5H:1V. Vegetation, topsoil and unsuitable materials will be removed from the Stage 1 WMF footprint to prepare the area for construction and waste placement, and will be stockpiled for use during future reclamation and closure phases. A geosynthetic lining system will be installed on the prepared Stage 1 foundation to minimize seepage from entering the underlying foundation soils. The geosynthetic lining system will include a 150 mm thick bedding layer to provide a suitable subgrade for the geosynthetics. The bedding layer will be covered by non-woven geotextile and 80 mil HDPE textured (both sides) geomembrane. A 300 mm thick drainage/cushion layer will be installed over the geomembrane to convey collected seepage from the WMF to the WMP.

Waste rock, coarse rejects and filtered tailings will be placed and compacted on the geosynthetic lining system at the same time to maintain the same elevation along the WMF crest. Placement will commence at the lowest elevation with a minimum offset distance from the WMF toe to the creeks set at approximately 30 m, and advance upslope. This approach will prevent ponding of water on the WMF surface and allow any runoff to shed from the WMF. During the winter months, snow will be removed from the interim surfaces as the material is placed. The following two zones will be constructed:

Perimeter Zone: This zone will be comprised primarily with waste rock with minor quantities of coarse rejects to form a 40 m wide, erosion resistant shell for the WMF. A 40 m wide, 1 m thick transition zone will be placed above the underdrain layer, prior to Perimeter Zone placement, for added liner protection.

Interior Zone: This zone will be comprised of co-disposed coarse rejects, waste rock and filtered dry-stack tailings. In some cases, the coarse rejects will be used to construct access roads within this zone to facilitate equipment routing and provide for dust mitigation during filtered tailings placement.

Bedding, underdrain and transition zone materials for construction will be sourced from underground mining development and from local excavations within the WMF basin. Processing will likely be required to produce the materials for the geosynthetic lining system.

The final WMF will be approximately 80 m in height with the crest constructed to El. 860 m and an overall outer slope angle of 3.5H:1V. The geosynthetic lining system and underdrainage layer will be extended periodically during operations as the footprint is expanded. Waste from

Years 3 through 12 will then be placed on the entire WMF footprint following the placement methods described above.

18.15.4 Water Management

The water management measures for the WMF are briefly described below. The approximate locations for the water management measures are shown in Figure 18.1.

Seepage Control

Geosynthetic Lining System – The lining system will be installed on the prepared basin subgrade to minimize seepage into the foundation. The geosynthetic lining system will include a granular bedding layer, non-woven geotextile cushion layer and 80 mil HDPE geomembrane.

Seepage Collection System – The subgrade and the geosynthetic lining system/underdrain layer will be graded and installed to route collected seepage to a seepage collection sump located at the northwest toe of the WMF. The collected seepage will be conveyed to the WMP and/or water treatment facility, treated as required and discharged to the environment. The seepage collection sump will be constructed of pre-cast concrete and will include inlet pipes extending a short distance upslope from the sump, a pump and outlet pipeline for water treatment and/or discharge.

Stormwater Management

Water Management Pond (WMP) – The WMP will be constructed to the northwest and downstream of the WMF to collect runoff from the two collection ditches. The WMP will provide temporary storage for storms up to and including the 1 in 10 year, 24-hour storm event (British Columbia Ministry of Environment; BCMOE, 2015). Runoff from storms between the 1 in 10 year, 24-hour storm event and the 1 in 200 year, 24-hour storm event will be conveyed via the WMP spillway to the environment (BCMOE, 2015). A floating pump and pipeline will be installed at the WMP to draw down the pond in a controlled manner following a storm event. The collected water will be discharged to the environment once the water quality meets local environmental objectives.

Diversion/Collection Ditches – Two diversion ditches are included in the concept to route noncontact water from the slopes above the WMF to the creeks. Two collection ditches are included in the concept to collect runoff (contact water) from the WMF outer slopes and convey the collected contact water to the WMP. The ditches will convey runoff from storm events up to and including the 1 in 200 year, 24-hour storm, as per local guidelines (BCMOE, 2015). Collection/diversion berms will be included in the ditch arrangements to provide the required conveyance capacities. This approach was adopted based on the assumed thin overburden profile at the site and the availability of clean, durable waste rock. The diches/berms will be approximately 1 m in depth/height with a 1 m base width and 2H:1V side slopes.

Water Treatment and Discharge – Seepage collected from the WMF underdrain may need to be transferred to a water treatment facility, treated as required and discharged to the environment. If the collected seepage meets local water quality objectives, the collected water will be transferred to the WMP. The design of a water treatment facility is being completed by Others.

Instrumentation

Instrumentation including vibrating wire piezometers and surface movement monuments will be installed at the WMF to monitor the performance of the facility.

18.15.5 Reclamation and Closure

The reclamation and closure concept for the WMF includes for the following.

The downstream slopes of the WMF will be progressively covered with soil and vegetated as the WMF is raised. A final soil/vegetative cover will be installed along the WMF crest at closure. Re-grading of the downstream slope is not anticipated to be required at closure. The WMF arrangement has been laid out to mimic the local topography and the closed facility will create a landform that closely resembles the surrounding landscape.

Ditches, WMP, Pipework and Appurtenances – These items will be decommissioned and removed as required. Disturbed areas will be re-graded and re-vegetated.

18.15.6 Material, Quantities and Cost Estimates

The schedules of materials and quantities, as well as initial and sustaining capital cost estimates for the WMF and water management measures are provided on Tables 2 and 3, respectively. The schedules of materials and quantities and cost estimate for WMF closure are provided on Table 4. The estimates were based on the conceptual level arrangements and were developed to an accuracy of approximately +/-50%. The following main items were not included in the cost estimates, as it is assumed that they will be designed/costed by others:

Tailings filtration plant construction and operation
Water treatment facility construction and operation
Access/haul road construction and maintenance

Unit rates were developed by KP using data from recent projects in the Yukon (Micon, 2020), local expertise in the Revelstoke region (Rokmaster, 2020) and recent project experience to assist with the preparation of the cost estimates. Lump sum allowances were included for select items based on recent and relevant experience. A 25% contingency has been included in the total capital cost and closure cost estimates. It is recommended that the total cost including this contingency be carried forward in the cost model for the Project. The costs are in 2020 CDN$ and no Net Present Value, inflation or discounting were applied to the costs.

The estimated costs for the WMF and water management measures are summarized below.

Initial Capital: The capital cost to construct the WMF and water management measures for two years of operations (Years 1 and 2) is estimated to be approximately $26.2 Million. It is assumed that the construction would be completed in Year 0, prior to plant commissioning. Sustaining Capital: The capital cost to construct the WMF and water management measures for ten years of operations (Years 3 through 12) is estimated to be approximately $67.4 Million (approximately $6.7 Million/year, evenly distributed over ten years).

Closure: The cost to implement the progressive and final closure measures is estimated to be approximately $2.3 Million. It is assumed that monitoring and construction will occur over a five-year period (Years 13 through 17, evenly distributed over five years). The closure period will be related to the duration that water from the seepage collection system may need to be transferred to the water treatment facility following the end of operations.

18.15.7 Recommendations/Potential Opportunities

The following key recommendations and potential opportunities are provided to assist Rokmaster with advancing the design of the WMF as part of future studies:

This study has assumed that the geotechnical conditions in the vicinity WMF are conducive to physical stability and providing appropriate borrow materials for processing and developing the proposed WMF concept. Site investigations should be completed including geotechnical drilling, test pit excavations, in situ testing, sampling, instrumentation installation and laboratory testing to gain an understanding of the geotechnical, hydrogeological, geological and geochemical site conditions and develop design parameters to support more detailed levels of design.
Physical and geochemical characterization of the tailings, coarse rejects and waste rock (including filtration test work on the tailings) should be undertaken to estimate in situ placed densities, stable slope angles, filtered tailings management requirements and potential ARD and Metal Leaching potential of the wastes.
Meteorological and hydrological data should be collected to develop climate normals and extreme storm event estimates, as well as to estimate the water management requirements at the site.
A seismic hazard assessment for the site should be completed to estimate the magnitude of extreme earthquake events.

19.0 MARKET STUDIES AND CONTRACTS

There are no definitive off-take contracts or other significant contracts in place for the project.

Market studies carried out by Rokmaster and its consultants were reviewed by Micon and are reflected in the estimated off-take terms for lead and zinc concentrates used in net revenue calculations for the base case study. Similarly, in the alternative case, suggested terms for thirdparty off-take of auriferous sulphide concentrates were also applied.

20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

The following sections provide a summary of environmental, permitting, social and community studies, and other potential issues for the project.

20.1 ENVIRONMENTAL STUDIES AND POTENTIAL ISSUES

Environmental information on the project is available from historical studies conducted by previous property owners in the 1980s, 1990s, and 2000s. Most of the available environmental information was summarized in the previous NI 43-101 Technical Report – Preliminary Economic Assessment (PEA) (Micon, 2012). Based on our review of the information provided as part of this study, it appears that little or only annual water sampling data have been collected between 2012 and late 2020, when weekly sampling commenced. A summary of the existing environmental information is presented below based on previous reports for the property. An extensive baseline data collection program covering atmospheric, aquatics, terrestrial and human environment components will be required to support the project environmental impact assessment (EIA), permitting, project design, and impacts management.

20.1.1 Overview

Elevations in the area of the project range from 700 to 3,050 metres above mean sea level. The topography is characteristic of the Selkirk Mountains.

The main watercourse on the property is Carnes Creek, which transects the area. Its source is the Durrand Glacier, which is east of the Property. McKinnon Creek is a tributary of Carnes Creek. The area surrounding the intersection of McKinnon and Carnes Creeks has been the focus of the majority of the work over the life of the mineral Property.

20.1.2 Climate and Meteorology

Some information on climatic conditions is available from previous reports. The summer weather is considered moderate with average temperatures between 16° to 30° Celsius, with long stretches of sun and rain. The average precipitation is 65 cm/year. Winters are long and are characterized by heavy snowfalls (1 to 4 metres) with cool temperatures (-15°C to +5°C). Snowfall typically occurs between October and May at higher elevations and between November and April at lower elevations. No local meteorological data are available for the project area.

20.1.3 Surface Water Quantity and Quality

The project is located at the headwaters of two creeks, namely McKinnon and Carnes creeks. Local hydrologic conditions for the two creeks were summarized in a 1990 report (Equinox, 1990). Reported stream levels and flows were highest in August, but snowmelt season was not measured. Additional baseline information on stream flows and water levels in the creeks may be available as a result of historical studies completed by provincial or federal agencies.

Some historic water quality information for the two creeks is available as a result of studies completed in the mid 1980s and the 2000s (Micon, 2012). Based on the available data, it appears that water quality in Carnes Creek is typical of glacial fed streams with moderate hardness (average hardness 118 mg/L CaCO3), an average pH of 8.0, with total dissolved solids and total suspended solids concentrations averaging 80 mg/L and 35 mg/L, respectively. Total iron and aluminum occasionally exceeded the BC freshwater quality guidelines for the protection of aquatic life. Water quality sampling resumed in late 2020.

20.1.4 Groundwater Quality and Quantity

There is currently no recent, comprehensive groundwater quality and quantity information for the project area. A groundwater baseline data collection program will be required to provide data for use in characterizing groundwater quality and quantities, as well as for use in the assessment of future impacts associated with the operations and closure of facilities such as the WMF.

20.1.5 Fisheries and Fish Habitat

Some limited information on fisheries and fish habitat in McKinnon and Carnes Creeks is available as a result of previous studies (Micon, 2012). Based on the existing information, kokanee salmon and bull trout were documented to occur in the creeks. No barriers to fish movement were found in the two creeks, although fish habitat appeared to be of poor quality perhaps due to the high levels of turbidity and low temperatures found in these glacier-fed creeks.

20.1.6 Vegetation, Forestry and Wetlands

Some vegetation information in the area is available as a result of surveys completed in the early to mid 2000s (Micon, 2012). The project site is located within the Interior Cedar-Hemlock Biogeoclimatic Zone (ICH) and within the very wet, cool variant subzone (ICHvk1). The area consists of areas of old growth western red cedar and western hemlock and areas of young planted spruce and Douglas fir with natural regeneration of red cedar, spruce and hemlock. Upper areas of the watershed, above 1,500 m, include the Engelmann spruce, subalpine fir (ESSF) biogeoclimatic zone and the Alpine Tundra (AT) zone at the peaks. Thirteen invasive plant species were noted during vegetation surveys completed in 2008. However, no rare plants were observed during the previous surveys.

20.1.7 Wildlife and Wildlife Habitat

Based on results of previous wildlife and wildlife habitat studies completed in the project area during the late 2000s (Micon, 2012), a number of wildlife species including red squirrel, mouse, chipmunk, marten, beaver, coyote, moose, and white-tailed and mule deer are known to occur in the area. Other wildlife species that are known to occur in the area include snowshoe hare, and bear. Although not detected during previous surveys, there is a potential for fisher, caribou, and wolverine to occur in the area.

Thirty-one species of birds were also recorded during a June 2008 breeding bird survey in the area (Micon, 2012). The most common species included the dark-eyed junco, MacGillivray's warbler, winter wren, and Swainson's thrush. Other species of note included nesting barred owls.

The western toad, which is federally listed as a species of Special Concern, was also observed during the June, 2008 survey.

20.2 WASTE MANAGEMENT

20.2.1 Waste Management

The main waste management issue for the project is the prevention and control of potential metal leaching/acid rock drainage (ML/ARD) from the WMF, and any acid generating or potentially acid generating (PAG) waste rock that is produced during mine development or operations. Non acid-generating (NAG) waste rock will be utilized for WMF construction. Forty-two percent of the total waste rock will be backfilled in underground voids where permitted. PAG waste rock, if encountered, will be deposited in the WMF and effluent managed to reduce any environmental impacts. Tailings will be deposited in the WMF and in underground voids as paste backfill. Waste rock and tailings management is further detailed in Section 18.15.

Some data on waste rock characterization and management are available from previous reports. Based on information summarized in the previous PEA (Micon, 2012), it appears that no evidence of existing acid generating rock could be detected from the 832 adit since development in 1991, indicating a substantial buffering potential exists in the deposit host rocks. The available information also suggests that development rock with greater than 0.6% sulphur may have uncertain acid generating potential. Arsenic was identified as a possible metal of environmental concern from the 832-level adit drainage water quality analyses.

Based on information contained in previous reports, the site was put on care and maintenance status in 2008 and the existing PAG waste rock pile was covered with overlapping panels of heavy-duty non-rip nylon tarp as a temporary measure to minimize infiltration into the pile and thereby minimize possible oxidation effects. In 2011, a lined pad for PAG waste rock was constructed at the site. In the spring of 2012, the existing PAG waste rock was re-located onto the lined pad.

Environmental requirements will guide construction decisions for management of tailings and waste rock produced during mining operations. Seepage and runoff from the tailings and waste rock management facility (WMF) will likely need to meet British Columbia's contaminated site criteria for groundwater, which correspond to general water quality standards (WQS) (BC, 2017), with an allowance for 1:10 dilution. Based on the results of kinetic testing work completed in 2011, leachate from the waste rock exhibited elevated sulphate, antimony, arsenic, cadmium, cobalt, and zinc. WMF seepage quality is expected to be similar to waste rock barrel draining and 830 portal seepage reported in the 2019 Reclamation Plan. Based on comparison with applicable WQS, the WMF is expected to require a liner and is conceptually designed and costed with a liner to comply with WQSs for arsenic, lead, and zinc, with more concentrated runoff requiring treatment for additional metals and, at times, sulphate. Seepage will be routed to the Wastewater Treatment System (WWTS) for treatment.

20.3 WATER MANAGEMENT

As noted previously, the project site is traversed by McKinnon and Carnes creeks. Historical water quality data are available from previous studies conducted at the project site. These data show that concentrations of most water quality parameters in the creeks are currently below the BC freshwater guidelines for the protection of aquatic life, with a few exceptions (total iron and total aluminum).

The current discharge from the 82 portal is permitted as required by the BC Environmental Management Act (EMA). BC has a requirement that discharge limits consider best available technology (BAT) evaluations and best management practices as part of decision-making. Discharge limits are required to meet the BC water quality guidelines for aquatic life. However, where this is not feasible, BC does allow the use of science-based environmental benchmarks (SBEBs) in setting discharge limits. SBEBs are considered only after BAT and best management practices have been considered and incorporated into development plans. SBEBs must be protective of the most sensitive aquatic species and life stages at a site.

Mining effluent discharges to fish-bearing creeks would also require approval under the federal Metal and Diamond Mining Effluent Regulations (MDMER) under the Fisheries Act.

20.3.1 Water Balance

Site water management will be a key issue for project design and permitting, including mine water that has been in contact with potential sources of contamination, seepage from the WMF, process water, and mine dewatering. Surface waters will be directed away from contact of surface facilities as required.

Strategies for water management will include:

• Keeping non-contact water runoff separate from mine water by diverting surface water runoff from precipitation away from active mine site areas,

Providing for the collection of surface water from disturbed areas to manage surface water erosion,
Recycling mine-contact water whenever possible.
Treatment as required to meet discharge standards.

An emergency response spill contingency plan will be developed to help prevent and respond to accidental release of contaminants as is standard practice for mine operations.

Water requiring treatment is assumed to be the excess water from mine dewatering activities, from the WMF, and surface water runoff from the site. The minimum flow of water that would need to be managed at the site is:

Non-contact stormwater:
- o Runoff from plant, roads, and other non-contact surfaces (assumed to be approximately 30 ha, with runoff coefficient of 40%).
Mine water:
- o Precipitation that falls on the WMF (assumed to be approximately 30 ha).
- o Mine dewatering to lower water table for new workings (assuming the volume of water removed is twice the volume to account for a small cone of depression and that excavation of mine workings is evenly spaced in time during a given year).
- o Mine dewatering to remove groundwater from existing workings (assuming existing excavations deliver about 270 m3 /day (3" pipe) and future excavations will be ten times existing, with no flooding of workings during operation). Backfilled workings are assumed to be filled in within a year of mining.
- o Mine dewatering to remove infiltration (assuming area draining to mine is half of mine footprint of 69 ha, with infiltration coefficient of 5%).

Mine dewatering flow estimates assumed fairly intact surrounding rock with a minimal cone of groundwater depression. If the rock walls of the mine workings stopes are significantly fractured, a much larger dewatering volume could be required to keep the mine dry, and dewatering flows could be up to an order of magnitude higher.

This water balance assumes about 20 m3 /day of net losses to evaporation and dust suppression activities. In reality, these losses will be higher during summer months when dust suppression is underway and temperatures support higher evaporation rates.

The estimated average and max monthly flows per year are listed in Table 20.1. A conceptual project water balance, with estimated average flows for Year 5 of operation is included as Figure 20.1.

Flows in m3/day	Year 1(average)	Year 5(average)	Year 5(peak)	Year 10(average)	Year 16(average)	Closure(average)
Non-contact runoff	331	331	879	331	331	0
Precipitation on tailings/waste management facility	827	827	2199	827	827	0
Dewatering for new workings	673	518	1037	326	244	0
Dewatering to keep existing workings dry	228	175	175	110	83	0
Infiltration from precipitation on overlying land	680	680	2464	680	680	0
Water to backfill past workings	-295	-295	-295	-295	-283	0
Water stored long-term in tailings	-375	-414	-414	-576	-742	109

Table 20.1 Estimated Major Sources and Sinks of Water

Figure 20.1 Conceptual Project Water Balance (with estimated average flows for Year 5 of operation)

The processing plant is expected to use about 700 m3 /day of water, which is similar to the amount of mine dewatering water expected. As such, water is not expected to be recycled through the plant. Process water will be sourced from mine dewatering water, and excess process water will be routed to treatment. Precipitation to the WMF will be collected in the water management pond and as seepage, and will also be available as process water.

20.3.2 Water Treatment Conceptual Basis

20.3.2.1 Water Treatment Targets

Water treatment targets were assumed to match BC water quality standards for freshwater aquatic life (BC, 2019), with values dependent on hardness and pH calculated based on monitoring data from the creeks. The 2019 Reclamation Report lists hardness concentrations in McKinnon and Carnes Creeks between 50 and 120 mg/L as CaCO3 and pH values between 7.9 and 8.2 (Haukan, 2019).

20.3.2.2 Mine Water Flows

Based on this initial evaluation, a water treatment system may have to treat up to 5,500 m3 /day through nanofiltration and 1,700 m3 /day through chemical precipitation during operations (peaking in Years 5 through 7) (Table 20.1). During closure, the only source of water requiring treatment is expected to be seepage from the WMF sourced from water within placed tailings pore spaces. This flow was assumed to be approximately 110 m3 /day, based on experience with similar projects. Seepage would be treated in a constructed wetland area located to the Northwest of the planned WMF, between the dam toe and the confluence of the creeks.

20.3.2.3 Mine Water Quality

All water is assumed to be captured on the site and used in plant processes. Mineral mass reporting to process water and ultimately requiring treatment is expected to be sourced primarily from mine dewatering and leaching from plant processes.

According to the 2012 PEA, rock proximal to the main orebody is expected to be similar to reported waste rock mineralogy (limestone, phyllite, and quartzite). Mine water quality coming out of the processing plant is assumed to be similar to leachate from waste rock barrels and portals reported in the 2012 Reclamation Plan. This water was generally not acidic, with pH values near 8.

Parameters assumed to require removal from mine water flows are:

Cationic metals: Cd, Co, Cu, Pb, Zn (up to 95% removal, limited by Zn).
Anionic metals: As, Se, Sb (up to 95% removal, limited by As).
Others: sulphate (up to 75% removal), mercury (up to 95% removal).

20.3.3 Water Treatment Conceptual Design

20.3.3.1 Non-Contact Stormwater Management during Mine Operations

A stormwater holding pond will be constructed to the northwest of the WMF, with the purpose of reducing solids in non-contact runoff from 30 ha prior to creek discharge. The pond is expected to occupy about one hectare, and to be lined with a low permeability soil to facilitate

conversion to constructed wetlands after closure. Non-contact runoff will be routed from the plant site via piping placed on the north side of the tailings and waste rock storage facility. Actual pond size will depend on the actual working area requiring stormwater collection and degree of treatment required.

20.3.3.2 Mine Water Treatment Processes during Mine Operations

Mine Water Treatment would have to remove cationic metals (up to 95%), anionic metals (95%) and sulphate (75%). The conceptual treatment process selected to remove this suite of constituents would consist of a two-stage high-density sludge metals precipitation with two separate steps:

Initial step to add iron and decrease pH (3-5) to support adsorption of arsenic and other anionic metals onto ferric oxide surfaces (ferrihydrite adsorption).
Final step to add lime up to pH 10 to precipitate cationic metals as metal hydroxides and sulphate as gypsum (hydroxide precipitation).

Chemical precipitation will be preceded by membrane separation using nanofiltration membranes for dilute flows requiring treatment, which serves to concentrate salts for more efficient chemical precipitation (of membrane concentrate) and to remove metals prior to discharge (of membrane permeate). A conceptual process flow diagram for water treatment sources and process flows is included as Figure 20.1. These technologies are all conventional technologies that have been demonstrated to treat these constituents in full-scale operations at similar mining operations. Alternate treatment processes that could also potentially be used to remove these constituents could include ion exchange or sorption with proprietary materials. As with chemical precipitation, these other treatment technologies may require separate processes for removal of anions, cations, and sulphur. A comprehensive trade-off study evaluating relative costs of different wastewater treatment technologies and process options is recommended prior to detailed design project phases.

Ferrihydrite adsorption operates by providing ferrihydrite surfaces to sorb anionic metal ions. This process will include a recycle sludge blend tank, two reactor tanks, and one clarifier. Solids from the low pH treatment steps will be removed using a clarifier. A portion of this sludge will be recycled back to the blend tank to maintain ferrihydrite surface area in the system, with the remaining solids removed from the water and managed as described below. The two reactor tanks will be operated at two separate pH values: one at pH 3 to target selenium and antimony sorption, and one at pH 6.5 to target arsenic sorption. Ferric sulphate may be added during this step to increase the concentration of iron solids.

Hydroxide precipitation will be operated by adding lime to raise solution pH and precipitate cationic metals as metal hydroxide solids. The process will operate at pH 9 to optimize removal of lead and zinc. This treatment process is also expected to remove some copper, aluminum, and other cationic metals. If additional metal removal is needed to meet water quality standards, a sulphide-based metal scavenger would be added at this phase. Mercury removal is not typically achieved through hydroxide precipitation, and would also benefit from

scavenger addition at this step, if necessary. A portion of the settled sludge will be recycled back to the blend tank to support a high-density sludge process.

Collected chemical precipitation sludge would be routed to a storage tank for mixing into mine backfill material. Future work should include simulation and testing of solid stabilization to determine the degree to which metals can be stabilized.

20.3.3.3 Equalization/Flow Balancing of Mine Water

Snowmelt has the potential to cause significant seasonal peaking factors in water flows. Equalizing snowmelt flows over a two-month duration is expected to require 50,000- 100,000 m3 of volume to equalize snowmelt from a wet year. Given the planned WMF pond size of 250,000 m3 (including freeboard), the WMF pond will be operated at a lower volume during the winter to make space for some equalization of spring snowmelt.

20.3.3.4 Conceptual WWTS Location and Equipment Sizing

The WWTS is expected to require approximately 350 m2 of space, and would be co-located with the processing plant, if possible, to minimize pumping and transportation costs.

The proposed treatment scheme is expected to require the following major equipment items:

Non-contact stormwater basin (1).
Nanofiltration membrane skids, with chemical feed and CIP skids (3).
Reactor feed pumps (2).
Ferrihydrite sludge/reaction tanks (3).
Ferrihydrite clarifiers (2).
Hydroxide sludge/reaction tanks (2).
Hydroxide clarifiers (2).
Carbon dioxide chemical storage and feed equipment (2).
Lime chemical storage and feed equipment (1).
Sludge recycle pumps (2).
Sludge storage tanks (2).
Blended permeate tank (1).

20.3.3.5 Closure-Phase Treatment

Mechanical water treatment is expected to be required for several years after the end of mining operations while closure activities are completed.

Seepage water quality is expected to be similar to the waste rock test barrels described in the 2019 Reclamation Plan (Haukan, 2020). To treat this seepage, the stormwater pond will be converted into a constructed wetland that will be designed to retain metals and sulphate. After closure, any seepage collected from the WMF would be routed to the constructed wetland prior to stream discharge.

After closure, runoff from the site is assumed to not require treatment.

20.4 PERMITTING REQUIREMENTS

All mining projects in BC are subject to approvals under the BC Mines Act and must comply with requirements of the Health Safety and Reclamation Code for Mines in BC. The Ministry of Energy, Mines and Petroleum Resources is responsible for the regulation of mines and mining activity, including mine health and safety aspects, closure and reclamation. Applications for mine permits, including exploration programs, are referred to other provincial agencies, including the Ministry of Environment & Climate Change, and to First Nations. Comments provided by other agencies are considered by the Chief Inspector of Mines when making a decision on whether or not to issue a permit.

Mining projects in BC may also be subject to various federal regulations depending on various factors, including size, potential for impacts on fisheries, species at risk, migratory birds and indigenous community concerns.

A number of mining and exploration permits have been issued for this project by BC regulatory agencies in the past. Permits currently in place include a general mine site permit (Permit MX-4-500) updated on August 31, 2021, by the BC Mines Branch for underground access, PAG storage, and underground diamond drilling and an underground wastewater discharge authorization #110409 issued on September 14, 2020 by the BC Ministry of Environment and Climate Change Strategy.

Some of the permits issued in the past dealt with specific activities such as explosives storage and use and forest road use.

Major mines in BC require many authorizations from many different federal and provincial agencies. The number and type of authorizations required varies from mining project to mining project, but typically include three primary provincial authorizations:

Environmental assessment (EA) certificates, issued under the BC Environmental Assessment Act (Environmental Assessment Office).
Permits issued under the Mines Act (Ministry of Energy, Mines and Petroleum Resources; BC EMPR).
Air and water discharge permits issued under the Environmental Management Act and its regulations (Ministry of Environment and Climate Change Strategy).

Additional permits and approval may be required, including the following:

Department of Fisheries and Oceans (DFO) approval.
Ministry of Environment and Climate Change Strategy approval.
Explosives Magazine Storage and Use Permit from the B.C. Ministry of Energy, Mines and Low Carbon Innovation.
Mining Lease from the B.C. Ministry of Forests, Lands, Natural Resource Operations and Rural Development.
Licence to Cut or Free Use Permit from the B.C. Ministry of Forests, Lands, Natural Resource Operations and Rural Development.
Navigable Waters approval.

In addition, permits and authorizations may also be required under other provincial acts and regulations.

On December 16, 2019, the new BC Environmental Assessment Act (2018) came into force. While projects with an Environmental Assessment already underway will continue under the old Act (2002), any new projects registered with the EAO after December 16, 2019 will undergo an environmental assessment under the new Act (2018) process. Some of the key changes included in the new BC Environmental Assessment Act (2018) relative to the previous requirements include the need for early engagement and use of the EIA process to advance reconciliation with indigenous communities.

The proposed project constitutes a reviewable project pursuant to the Reviewable Projects Regulation (B.C. Reg. 243/2019) of the BC Environmental Impact Assessment Act (2018) because the proposed Project would have a production capacity of greater than or equal to 75,000 tonnes per year of mineral ore.

The Environmental Assessment Office (EAO) is the agency that manages the review of proposed major projects in BC, as required by the Environmental Assessment Act (EAA) and Reviewable Projects Regulation. The EAO ensures proposed major projects meet provincial environmental, economic and social objectives. The process evaluates proposed projects that are reviewable under the EAA for potential adverse environmental, economic, social, heritage and health effects. If a project is approved, the EAO verifies and enforces compliance with the conditions set out in environmental assessment certificates.

Although timelines for completion of EIAs vary by project complexity, issues identified and the level of interest or involvement of stakeholders, based on recent experience, EIAs studies, review and decision-making can take anywhere between two and five years. To expedite the regulatory review process, some of the other permits required by the project under other statutes (such as waste (air and water) discharge permits) are typically included as part of the integrated application during the EIA process.

For the purposes of the PEA, it is assumed that permitting will take approximately two to three years. Permitting costs are included as indirect costs and are applied in the pre-production period. Future studies to support environmental assessment should include:

Evaluating groundwater quality and quantity and running a groundwater protection model to predict groundwater impacts.
Additional geochemistry work to predict water quality.
- o Humidity cells (e.g., ASTM D5744)
Additional chemistry and ABA testing on a larger suite of samples of all rock types.
Pilot and bench tests to determine water treatment equipment requirements and operating conditions.

Mine reclamation costs in BC require financial assurance prior to construction.

The project is also subject to the requirements of the new (2019) Canadian federal Impact Assessment Act (Physical Activities Regulations) 1 . BC EAO and the federal Impact Assessment Agency of Canada have established a framework that allows for the implementation of the principle of 'one project, one assessment'. Under this arrangement, the parties can cooperate on EIA matters while exercise their respective powers and duties. Such substituted impact assessments provide benefits to proponents, Indigenous peoples and the public by reducing workload and streamlining participation while ensuring that the expertise of both governments is applied.

Fisheries and Oceans Canada and Environment Canada are responsible for the Fisheries Act, and the project may be subject to requirements under this act if it results in impacts on fisheries and fish habitat fish-bearing local creeks. In addition, the project is likely to require a permit from Natural Resources Canada under the Explosives Act. Other federally managed legislation may also apply to the project, such as Migratory Birds Convention Act, Species at Risk Act and the Navigation Protection Act which regulate impacts on migratory birds, endangered species and waterways, respectively.

For the purpose of the updated PEA, it is assumed that a substituted EIA process for the project will take approximately two to three years. This includes baseline data collection, indigenous and public consultations, impact assessment, development of mitigation and management measures, responses to SIRs, public hearings, and post-approvals permitting.

20.5 ENVIRONMENTAL, SOCIAL AND HEALTH MANAGEMENT SYSTEM

In line with best practices, it is expected that an Environmental, Social & Health Management System (ESHMS) will be available for use during the pre-construction, construction,

1 a) a new metal mine, other than a rare earth element mine, placer mine or uranium mine, with an ore production capacity of 5,000 t/day or more; b) in the case of an existing metal mine, other than a rare earth element mine, placer mine or uranium mine, if the expansion would result in an increase in the area of mining operations of 50% or more and the total ore production capacity would be 5,000 t/day or more after the expansion.

operations and eventual closure and decommissioning phases of the project. The ESHMS will be used to manage compliance, impacts management and monitoring and to maintain the social license to operate. It is expected that the ESHMS will be developed and implemented based on a Plan, Do, Check and Act framework as is typical of most management systems. Depending on the components and complexity of the project, the following specific management plans may be required:

Surface Erosion Prevention and Sediment Control Plan.
Soil Management Plan.
Construction Environmental Management Plan.
ML/ARD Management Plan.
Mine Site Water Management Plan.
Discharge Management Plan.
Vegetation Management Plan.
Invasive Plant Management Plan.
Wildlife Management Plan.
Archaeological Management and Impact Mitigation Plan.
Mine Emergency Response Plan.
Mine Site Traffic Control Plan.
Fuel Management and Spill Control Plan.
Dust Management Plan.
Chemicals and Materials Storage, Transfer, and Handling Plan.
Waste (Refuse and Emissions) Management Plan.

For the purposes of the PEA, it is assumed that an operation of this size will employ one Environmental, Social, Health and Safety (ESHS) Manager, one Indigenous Communities Advisor, one Environmental Advisor and one Health & Safety Advisor, with additional help coming from consultants and contractors as necessary. Annual environmental, social, health and safety operating costs are estimated at $1 million, in addition to costs associated with the four ESHS staff.

20.6 SOCIAL AND COMMUNITY ISSUES

The area north of Revelstoke is a sparsely populated and relatively undeveloped region of the province. Forestry, electric power generation & transmission, railway transportation, and recreational tourism are the main sources of income. Community and socio-economic impacts of the project can potentially be very favourable for the region, as new long-term opportunities are created for local and regional workers.

20.7 FIRST NATIONS

The project falls within the traditional territories of three indigenous populations, namely the Okanagan, Shuswap and Ktunaxa First Nations, each of which is made up of one or more bands (Table 20.2). Based on the previous PEA (Micon, 2012), consultations have taken place with the indigenous communities during previous exploration phases and are likely to continue with the goal of developing working relationships necessary as the project plans become more defined and the project moves closer to development.

Nation	Band
Ktunaxa Nation Council (KNC)	Akisq'nukFirst Nation in Windermere.
in Cranbrook.
	Shuswap Indian Band in Inveremere. (Note -referrals are sent to
Shuswap Nation Tribal Council	KinbasketGroup of Companies with a cc to Chief and Council -Shuswap
(SNTC) in Kamloops	Indian Band.)
	Splatsin First Nation in Enderby
	Little Shuswap Indian Band in Chase
	Adams Lake Indian Band in Chase
	Neskonlith Indian Band in Chase
Okanagan Nation Alliance	Okanagan Indian Band (OKIB) in Vernon.
(ONA) in Westbank.

Table 20.2 First Nations with Potential Interest in the Project Area

BC's new Environmental Impact Assessment Act (2018) includes new requirements for enhancing public confidence by ensuring First Nations, local communities and governments and the broader community can participate in the EIA process through a robust and transparent manner. A key requirement is early engagement and planning.

Additionally, one of the stated goals of the new EIA process is advancing reconciliation with indigenous communities. This includes supporting the BC government in implementing the United Nations Declaration on the Rights of Indigenous People (UNDRIP), which BC has recognized and has pledged to implement.

These new requirements will require focused attention during the next phases of the project. Early planning and active engagement with the EAO and the potentially impacted indigenous communities will be required to minimize risks and delays to project approvals and implementation.

Rokmaster will engage and collaborate with federal, provincial, regional, and municipal government agencies and representatives as required with respect to topics such as land and resource management, protected areas, official community plans, environmental and social baseline studies, and effects assessments. Rokmaster will consult with the public and relevant stakeholder groups, including land tenure holders, businesses, economic development organizations, businesses and contractors (e.g., suppliers and service providers), and special

interest groups (e.g., environmental, labour, social, health, and recreation groups), as appropriate.

20.8 CLOSURE AND RECLAMATION REQUIREMENTS

Under the BC Mines Act and Health, Safety and Reclamation Code for Mines in BC, companies are required to obtain a permit approving the mine plan, a program for protection of the land and watercourses, and a reclamation program before starting work. Annual reclamation reports must also be submitted to the BC EMPR. Companies are required to post reclamation bonds prior to the start of activities.

Based on a review of the existing project information, annual closure and reclamation plans were prepared and submitted to regulators, with the last such plan having been submitted in 2020. The estimated total costs of outstanding reclamation liabilities for the project based on Huakan (2020) was $72,500.

A number of closure and reclamation activities will need to be accomplished in the future once mining activities are completed:

Decommissioning and removal of the plant site and ancillary facilities.
Closure and securing underground works.
Closure and reclamation of the waste management facility.
Closure, recontouring and reclamation of waste rock dumps (if separate from WMF).
Stabilization, reclamation and revegetation of all disturbed including accesses roads.
Ongoing dam stability and reclamation.
Conversion to non-mechanical water treatment and decommissioning of mechanical treatment equipment.

A Closure and Reclamation Plan will be developed as part of the EIA and refined for the permitting process. In summary, the mine closure concept is to meet water quality objectives without ongoing treatment for ARD. Closure planning will include dialogue with First Nations and stakeholders to determine post-mining land use objectives and necessary investigations required to achieve and monitor those objectives.

21.0 CAPITAL AND OPERATING COSTS

Micon's assessment of the capital and operating costs for the project base case, comprising an underground mine, on-site concentrator and pressure-oxidation (POX) plants, described below. In the base case, lead and zinc concentrates are sold, while gold concentrates are treated onsite in the POX plant to recover gold into doré bars.

An alternative scenario is also considered, in which no POX plant is built, and gold-bearing concentrates are shipped for treatment by third parties.

The following cost estimates are expressed in fourth quarter 2020 Canadian dollars, without provision for escalation. Where appropriate, an exchange rate of US$0.77/CDN$ has been applied. The expected accuracy of the estimates is ±30%.

21.1 CAPITAL COSTS

The capital cost estimate for the base case includes the costs of developing and equipping the underground mine, concentrator and POX plants, tailings storage facility and other on-site infrastructure. It takes into account the existing adit, portal, assorted mobile equipment, surface workshops, and other infrastructure that is present at the site.

Total capital costs for the base case are forecast as shown in Table 21.1.

Area	Initial Capital($'000)	Sustaining Capital($'000)	LOM TotalCapital ($'000)
Mining	33,794	194,252	228,046
Processing Plant	250,690	7,840	258,170
Site Infrastructure	41,158	63,137	104,295
Owner's Costs	7,930	0	7,930
Contingency	62,392	8,950	71,342
Total	395,963	273,820	669,783
Closure Costs	0	6,500	6,500
Grand Total	395,963	280,320	676,283

Table 21.1 Capital Cost Summary – Base Case

An alternative development option assessed as part of the PEA is to develop the mine and concentrator without the POX plant. In this scenario, gold-bearing concentrates are sold to a third party for treatment and refining.

Total capital costs for the alternative, toll-treatment case are forecast as shown in Table 21.2.

Area	Initial Capital($'000)	Sustaining Capital($'000)	LOM TotalCapital ($'000)
Mining	33,794	194,252	228,046
Processing Plant	114,370	7,840	121,850
Site Infrastructure	36,368	63,137	99,505
Owner's Costs	7,930	0	7,930
Contingency	34,170	8,950	43,120
Total	226,631	273,820	500,451
Closure Costs	0	6,500	6,500
Grand Total	226,631	280,320	506,951

Table 21.2 Capital Cost Summary – Alternative Case

21.1.1 Mining Capital

Mining capital costs have been estimated on the basis of an owner-operated fleet of trackless mining equipment, developing conventional ramp haulage access (5 x 5 m) and lateral development headings (3.5 x 4 m).

The value of equipment already owned by Rokmaster is treated as a sunk cost and so is not reflected in these estimates or in the project cash flow. Micon estimates the replacement cost of that equipment to be approximately $4.0 million, comprising 2 - 6yd scoops, 2 - 30 tonne haul trucks, generator suitable for mine development / standby power, twin-boom jumbo, pumps, fans, switchgear, fully-equipped mechanical shop, office with communication, and assorted ancillary equipment.

The initial mining capital cost estimate set out in Table 21.3 is the same for both the base case (POX) and alternative scenario.

Area	Initial Capital($'000)	Sustaining Capital($'000)	LOM TotalCapital ($'000)
Mobile Equipment	4,616	89,384	94,000
Main Ventilation Fan	-	500	500
Ramps	3,667	25,667	29,333
Air Raises	1,709	5,128	6,838
Manways & Ore-passes	855	5,983	6,838
Horizontal Development(waste only)	22,947	67,591	90,538
Total Mining Capital	33,794	194,252	228,046

Table 21.3 Capital Cost Summary – Mining

21.1.2 Processing Capital

21.1.2.1 Direct Costs

Processing capital costs for the base case have been forecast on the basis of estimates of the installed cost for primary equipment, with appropriate factors applied for piping, civils (steel, concrete, earthworks), electrical, and instrumentation. The cost of other processing direct cost items such as the mill buildings and laboratory have been estimated separately.

The estimated capital costs for the pressure oxidation and gold recovery circuits were provided to Micon by Canenco.

The base case processing capital cost estimate is set out in Table 21.4 alongside the alternative case (without a hydrometallurgical plant to treat the gold-bearing concentrates). In each case, a provision for sustaining capital costs of $7.48 million is made in the cash flow model.

Area	Base Case($'000)	Alternative Case($'000)
Crushing and Pre-concentration	7,480	7,480
Grinding and Gravity Circuit Area	5,710	5,710
Flotation Areas	13,130	13,130
Concentrate Dewatering	3,870	3,870
Tailings Dewatering and Backfill Plant	2,570	2,570
Water Systems	1,240	1,240
Installation, concrete, steel, labouretc.	27,190	27,190
Electrical and Automation	12,820	12,820
Piping	7,770	7,770
Site preparationand Plant Buildings	14,500	9,500
Sub Total Mineral Processing Direct Costs	96,280	91,280-
POX/HAC/CCD/Neutralization	72,660	-
CIL	18,760	-
ADR/Electrowinning/Smelting	9,350	-
CN detox.	1,960	-
Reagents	2,150	-
Sub-Total Hydrometallurgical, Gold Recovery	104,880	-
Total Processing Direct CapitalCosts	201,160	91,280

Table 21.4 Initial Capital Cost Summary – Processing

21.1.2.2 Indirect Costs

Indirect costs, including EPCM services, construction indirect costs, commissioning, freight, spares, first fills and insurance charges are based on factors applied to the direct cost estimate.

The initial indirect capital cost estimate for the base case and the alternative case are set out in Table 21.5. A contingency of 20% has then been added to both the direct and indirect estimates.

Area	Base Case($'000)	AlternativeCase ($'000)
EPCM services	30,150	14,420
Freight and spares (for mech. equip.)	3,930	1,940
First fill reagents (2 months supply)	3,430	960
Construction indirects	8,040	3,840
Commissioning incl. vendor reps.	1,970	970
Insurance	2,010	960
Total Plant Indirect CapitalCosts	49,530	23,090

Table 21.5 Initial Capital Cost Summary – Indirect

21.1.3 Infrastructural Capital

21.1.3.1 Initial Infrastructural Capital

The base case infrastructural capital cost estimate set out in Table 21.6.

Area	Base Case($'000)	Alternative Case($'000)
Access Road upgrade	224	224
Site Roads	90	90
Waste Management Facility	16,140	11,440
Power Line	11,500	11,500
Power Distribution	7,500	7,500
Emergency Power	450	360
Fuel Storage & Distribution	150	150
Water Supply	650	650
Info & Communications Systems	134	134
Mobile Equipment -Surface	450	450
Warehouse, inventory control system	1,250	1,250
Assay Laboratory, incl. instruments	1,120	1,120
Reagent Storage	1,500	1,500
Total Infrastructural Capital	41,158	36,368

Table 21.6 Initial Capital Cost Summary – Infrastructure

Infrastructural capital costs for the base case include the first phase of construction on a waste storage facility for tailings or waste development rock that cannot be used as backfill or otherwise stowed underground.

In addition, a power supply to the site, upgrading of the access road, provision of a supply of fresh water and a wastewater treatment plant are recognised as infrastructural requirements that are common to both the base case and the alternative scenario.

21.1.3.2 Sustaining Infrastructural Capital

Sustaining infrastructural capital comprises a provision of $72.09 million for expansion of the dry-stack tailings and waste rock management facility (WMF), to be expended progressively over the LOM period. This amount includes a contingency of 20%.

21.1.4 Owner's Costs

21.1.4.1 Initial Capital

The initial capital cost estimate for the base case and the alternative case includes a provision for owner's cost as set out in Table 21.7.

Area	Base Case($'000)	AlternativeCase ($'000)
Staffing, recruitment and training	2,400	2,400
Owners management construction team	1,350	1,350
Owner's preproduction manpower	1,380	1,380
Permitting, environmental monitoring	2,800	2,800
Total Indirect Capital	7,930	7,930

Table 21.7 Initial Capital Cost Summary – Owner's Costs

21.1.4.2 Sustaining Capital

Owner's costs incurred during the operational phase of the project are included in General and Administration operating costs.

21.1.5 Mine Closure Costs

A provision has been made for pre-production bonding in the amount of $6.5 million to offset mine closure costs. A provision for post-closure monitoring of the facility is included in that amount.

21.1.6 Contingency

A contingency, amounting to 20% of the direct and indirect capital costs for processing and infrastructure, is included in the initial capital cost estimates, totalling $62.4 million for the base case and $34.2 million for the alternative case, respectively.

21.2 OPERATING COSTS

21.2.1.1 Base Case

Estimated LOM total cash costs for the base case are summarized in Table 21.8.

Area	Life-of-Mine Cost($ 000)	Unit Cost$/t milled	Unit CostUS$/oz Au Eq.
Mining	626,692	62.42	323.56
Processing	653,287	65.07	337.29
General & Administrative	76,202	7.59	39.34
Total Cash Costs	1,356,180	135.08	700.20

Table 21.8 LOM Total Cash Costs – Base Case

21.2.1.2 Alternative Case

Estimated LOM total cash costs for the alternative case are summarized in Table 21.9.

Area	Life-of-Mine Cost($ 000)	Unit Cost$/t milled	Unit CostUS$/oz Au Eq.
Mining	626,692	62.42	478.56
Processing	249,758	24.88	19.72
General & Administrative	76,202	7.59	58.19
Total Cash Costs	952,652	94.89	727.47

Table 21.9 LOM Total Cash Costs – Alternative Case

21.2.2 Mine Operating Costs

Mine operating costs are the same for the base case and the alternative.

All mine development and operating costs are based on an owner-operated fleet, with unit rates for drill, blast, load and haul applied to development metreage measured from the proposed underground layout. Estimates are based on Micon's analysis of the fleet and labour requirements to meet the demands of the development and production schedules, with provision for ancillary costs of mine technical services, supervision and equipment maintenance. The cost estimate also considers the volume of backfill material to be placed and haulage of excess waste material to surface.

Mine development costs will be capitalized, except for crosscuts, sills and other short-lived development that services only one group of stopes.

Table 21.10 summarises the mining operating cost estimate.

Area	Unit	Unit Cost($/m)	Quantity	Life-of-MineCost ($ 000)	Cost($/t milled)
Ramps	m	5,500	5,333	29,333	2.92
Air Raises	m	3,500	1,954	6,838	0.68
Manways & Ore-passes	m	3,500	1,954	6,838	0.68

Table 21.10 Mine Operating Costs (Base Case)

Area	Unit	Unit Cost($/m)	Quantity	Life-of-MineCost ($ 000)	Cost($/t milled)
Horizontal Development (waste)	m	4,500	32,712	90,538	9.02
Other Horizontal Development	m	4,000	32,712	50,370	5.02
Stopes	000't	47.50	7,741	367,674	36.62
Sills	000't	73.33	2,845	208,648	20.78
S/Total Mining costs				760,238	75.72
Development costs capitalized				(133,546)	(13.30)
Total Mine Operating Costs				626,692	62.42

21.2.3 Processing Operating Costs

21.2.3.1 Base Case

The base case process operating cost estimate is based on labour, power, process consumables and equipment maintenance requirements.

Including the pressure oxidation circuit, power demand is provisionally estimated at 101.7 kWh/t. With a tariff of $0.06/kWh, power for processing accounts for an average of $6.10/t milled.

A total head count of 75 people will be required in the plant, including technical support and maintenance staff. Annual maintenance costs for equipment and buildings are based on factors applied to original cost of supply.

Provision is made in the operating cost estimate for crusher and mill liners, grinding media, lime, cyanide and CN-destruction reagents and gold-room supplies. It is assumed that limestone for neutralization will be sourced and ground locally.

Table 21.11 summarises the process operating cost estimate for the base case.

Area	Main Zone$/t milled	Yellow Jacket$/t milled	Life-of-MineCost ($ 000)	Average Cost$/t milled
Labour	6.24	6.24	62,674	6.24
Electrical power	2.92	2.10	28,795	2.87
Process consumables	13.70	11.41	136,049	13.55
Maintenance	2.24	1.81	22,240	2.22
Concentrate Treatment	42.96	-	403,529	40.19
Process Operating Costs	68.06	21.56	653,287	65.07

Table 21.11 Process Operating Costs (Base Case)

21.2.3.2 Alternative Case

Processing costs for the alternative case are based on unit costs for production and sale of concentrates only. No downstream treatment and refining would be carried out at the project.

Table 21.12 summarises the process operating cost estimate for the alternative case.

Area	Main Zone$/t milled	Yellow Jacket$/t milled	Life-of-MineCost ($ 000)	Average Cost$/t milled
Labour	6.24	6.24	62,674	6.24
Electrical power	2.92	2.10	28,795	2.87
Process consumables	13.70	11.41	136,049	13.55
Maintenance	2.24	1.81	22,240	2.22
Concentrate Treatment	-	-	-	-
Process Operating Costs	25.10	21.56	249,758	24.88

Table 21.12 Process Operating Costs (Alternative Case)

21.2.4 General and Administrative Costs

21.2.4.1 Base Case

Provision is made in the project economic evaluation for General and Administrative (G&A) costs of $5.04 million per year ($6.00/t), plus an annual allowance of $0.34 million for maintenance of the waste management facility. In addition, the interest portion of leasing costs for the mobile mining equipment fleet, totalling $11.87 over the LOM, bring the total G&A operating costs amount to $76.2 million, equating to a cost of $7.59/t milled.

Equipment lease costs assume a down-payment of 20% of the purchase price, the full balance paid off over 5 years with an annual interest rate of 6% in real terms, and no residual amount.

21.2.4.2 Alternative Case

Estimated G&A costs in the alternative case are the same as for the base case.

21.2.5 Selling Costs

The transport, treatment and refining costs of concentrates sold are included in the calculation of net smelter returns – see Section 22.0.

Selling costs for doré bars (base case only) comprise bullion transport, insurance and refining charges, and are provisionally estimated at US$5.00/oz gold and US$0.50/oz silver.

22.0 ECONOMIC ANALYSIS

22.1 CAUTIONARY STATEMENT

The results of the economic analyses discussed in this section represent forward-looking information as defined under Canadian securities law. The results depend on inputs that are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.

Information that is forward-looking includes:

Mineral Resource and Mineral Reserve estimates.
Assumed commodity prices and exchange rates.
The proposed mine production plan.
Projected mining and process recovery rates.
Assumptions as to mining dilution.
Capital and operating cost estimates and working capital requirements.
Assumptions as to closure costs and closure requirements.
Assumptions as to environmental, permitting and social considerations and risks.

Additional risks to the forward-looking information include:

Changes to costs of production from what is assumed.
Unrecognized environmental risks.
Unanticipated reclamation expenses.
Unexpected variations in quantity of mineralized material, grade or recovery rates.
Geotechnical or hydrogeological considerations differing from what was assumed.
Failure of mining methods to operate as anticipated.
Failure of plant, equipment or processes to operate as anticipated.
Changes to assumptions as to the availability and cost of electrical power and process reagents.
Ability to maintain the social licence to operate.
Accidents, labour disputes and other risks of the mining industry.
Changes to interest rates.
Changes to tax rates and availability of allowances for depreciation and amortization.

This preliminary economic assessment is preliminary in nature; it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the preliminary economic assessment will be realized.

22.2 BASIS OF EVALUATION

Micon has prepared its assessment of the Project on the basis of a discounted cash flow model, from which NPV can be determined. Assessments of NPV are generally accepted within the mining industry as representing the economic value of a project after allowing for the cost of capital invested.

The objective of the study was to determine the potential viability of an underground mine with (or alternatively, without) a processing plant on site. In order to do this, the cash flow arising from the base case and the alternative case have both been forecast, enabling a computation of the respective NPVs to be made. The sensitivity of these NPVs to changes in the base case assumptions are then examined.

22.3 MACRO-ECONOMIC ASSUMPTIONS

22.3.1 Exchange Rate and Inflation

All results are expressed in Canadian dollars except where stated otherwise. Metal prices and revenues are converted at the rate of CDN$1.30/USD. Cost estimates and other inputs to the cash flow model for the project have been prepared using constant, fourth quarter 2020 money terms, i.e., without provision for escalation or inflation.

22.3.2 Weighted Average Cost of Capital

In order to find the NPV of the cash flows forecast for the project, an appropriate discount factor must be applied which represents the weighted average cost of capital (WACC) imposed on the project by the capital markets. The cash flow projections used for the evaluation have been prepared on an all-equity basis, except insofar as the mining equipment is assumed to be leased. This being the case, WACC is equal to the market cost of equity, which in this study is calculated according to the capital asset pricing model (CAPM).

CAPM depends upon the risk-free rate of return, the market premium for equity and the value of beta (β) for the project. In recent years, the risk-free rate has fallen close to zero, while it has been assumed here that the risk premium for equity has remained constant at around 5%.

While gold properties may typically have a β of around 1.0, for base metal producers the value of β is often between 1.5 and 3.0. Here, the base case cash flow shows approximately 67% of net revenues are derived from gold and silver doré, with the remaining 33% from lead and zinc concentrates. Accordingly, Micon has assumed in its base case that the project could expect a

β value of 1.5. Using CAPM, this yields a real annual discount rate of 7.5% for the base case. Micon also tested the sensitivity of NPV over a range of real discount rates from 5% to 10%.

22.3.3 Expected Metal Prices

In the base case, project revenues will be generated from the sale of zinc and lead concentrates and of gold doré bars from the treatment of auriferous concentrates. In the alternative case, all three concentrates are sold for treatment elsewhere.

The Project has been evaluated using constant metal prices of US$1,561/oz Au, US$20.55/oz silver, US$0.91/lb lead and US$1.07/lb zinc. These prices reflect consensus average long-term price forecasts published by a major commercial bank at the end of October, 2020.

Figure 22.1 presents monthly average prices over the past ten years and shows the upward trend for gold and silver over the past two years, while lead and zinc have remained steady.

22.3.4 Taxation Regime

Canadian federal and British Columbia provincial income and mining taxes have been provided for in the economic evaluation.

22.3.5 Royalty

Micon understands that no royalties are payable on the Revel Ridge deposit.

22.4 TECHNICAL ASSUMPTIONS

The technical parameters, production forecasts and estimates described earlier in this report. are reflected in the base case cash flow model and are summarised below.

22.4.1 Product Offtake

Table 22.1 summarizes the terms for product off-take. Micon considers these generic terms to be representative of the market for similar products.

		Lead Conc	Zinc Conc	Base CaseGold doré	Alternative CaseGold Conc.
Payability	Gold	46%	43%	99.5%	65%
(LOM avg.)	Silver	65%	21%	98.0%	60%
	Lead	85%	-	-	-
	Zinc	-	85%	-	-
Transport costs	CDN$/wmt	187.00	100.00	Incl.	179.00
Treatment Charge	US$/dmt	n/a	220.0	n/a	n/a
Refining Charges	US$/oz Au		5.00	5.00
	US$/oz Ag		0.50	0.50
Penalties	US$/dmt	171.95	28.00

Table 22.1 Product Offtake Terms

22.4.2 Production Schedule

Figure 22.2 shows the annual tonnages of material milled, together with the overall gold production for the base case.

Figure 22.2 LOM Production Schedule

22.5 OPERATING MARGIN (BASE CASE)

Figure 22.3 shows the annual cash operating costs, compared to the net sales revenue, demonstrating that the project maintains a significant operating margin over the first five years, before the margin narrows (but remains positive) over the remainder of the LOM. Over the LOM, the operating margin is projected to average 55%.

Figure 22.3 LOM Net Revenue and Operating Costs (Base Case)

22.6 PROJECT CASH FLOW (BASE CASE)

The LOM base case project cash flow is presented in Table 22.2. Annual cash flows are set out in Table 22.3 and summarized in Figure 22.4.

	LOM Total$'000	$/t Milled	US$/ozAuEq
Net Revenue	3,014,454	300.25	1,556.36
Mining costs	626,692	62.42	323.56
Processing costs	653,287	65.07	337.29
General & Administrative costs	76,202	7.59	39.34
Total Cash Cost	1,356,180	135.08	700.20
Net cash operating margin	1,658,273	165.17	856.17
Initial capital	395,963	39.44	204.44
Sustaining capital	273,820	27.27	141.37
Closure bond	6,500	0.65	3.36
Net Cash flow before tax	957,562	97.81	507.00
Taxation	350,220	35.06	181.72
Net Cash flow after tax	607,342	62.75	325.28
All-in Sustaining Cost per ounce (AISC)			841.57
All-in Cost per ounce (AIC)			1,046.01

Table 22.2 Life-of-Mine Cash Flow Summary – Base Case

Table 22.3 Base Case Life of Mine Annual Cash Flow

Period	Units	LOMTotal	Yr-2	Yr-1	Yr.1	Yr.2	Yr.3	Yr.4	Yr.5	Yr.6	Yr.7	Yr.8	Yr.9	Yr.10	Yr.11	Yr.12
Tonnes milled (t'000)	t'000				839.5	839.5	840.0	839.5	839.5	839.5	839.5	839.5	839.5	839.5	839.5	804.9
Mill Head Grade -gold	g/t Au				6.48	6.10	6.10	5.90	5.49	3.46	3.01	2.95	2.67	2.41	2.07	0.85
Mill Head Grade -silver	g/t Ag				47.28	52.22	61.49	62.75	56.55	40.57	34.40	58.88	52.58	39.36	37.89	48.35
Mill Head Grade -lead	% Pb				1.49	1.82	2.09	2.15	1.80	1.45	1.27	1.55	1.58	1.36	1.43	1.76
Mill Head Grade -zinc	% Zn				2.43	3.05	3.31	3.15	2.90	2.70	2.33	2.26	2.42	3.02	3.59	4.12
Payable gold	koz Au	1,068.4			144.12	143.98	143.27	138.77	129.43	74.62	63.82	62.19	56.33	50.35	43.27	18.28
Payable silver	koz Ag	8,281.8			571.45	717.37	847.98	890.53	815.54	585.59	479.84	809.86	769.60	560.79	520.16	713.12
Payable lead	000'lb	254,919			16,062	22,884	26,431	27,684	24,035	19,528	16,835	19,417	20,354	18,103	18,505	25,081
Payable zinc	000'lb	449,822			25,237	36,699	40,699	39,518	36,627	33,984	29,815	28,284	29,784	39,848	46,825	62,503
Gold Equiv. (payable oz)	koz Au	1,490			168.08	179.24	183.69	179.93	167.06	106.21	91.00	93.62	88.31	83.53	79.55	69.68
Net Sales Revenue	$M	3,014.5			340.1	362.7	371.6	364.0	338.0	214.9	184.1	189.4	178.7	169.0	160.9	141.0

Mining	$M	626.7			49.7	50.8	50.4	51.3	52.9	52.1	52.0	53.4	55.4	54.0	53.8	50.9
Processing	$M	653.3			57.1	57.1	57.2	57.2	57.1	57.2	57.2	57.1	57.1	49.6	49.7	39.7
G&A	$M	76.1			5.9	6.8	6.4	6.1	6.8	6.4	6.0	6.8	6.4	6.0	6.8	5.9
Total Cash Costs	$M	1,356.2			112.7	114.7	114.0	114.6	116.8	115.7	115.2	117.3	118.9	109.6	110.3	96.5

Net cash operating margin	$M	1658.3			227.4	248.0	257.6	249.4	221.2	99.2	68.9	72.1	59.8	59.4	50.6	44.5

Initial capital	$M	396.0	114.2	281.8
Sustaining capital	$M	273.8			25.3	22.6	26.5	24.8	23.9	22.0	18.4	24.3	19.5	16.3	26.8	23.5
Closure provision	$M	6.5		6.5	-	-	-	-	-	-	-	-	-	-	-	-
Change in working capital	$M	-			22.1	4.5	1.5	0.0	(1.5)	(7.8)	(1.8)	1.4	0.2	(0.5)	0.8	(18.9)
Net Cash flow before tax	$M	982.0	(114.2)	(288.3)	180.0	220.9	229.6	224.6	198.8	85.1	52.3	46.5	40.1	43.5	23.0	39.9
Taxation	$M	352.0			4.6	50.1	59.0	72.4	67.2	24.2	15.0	16.1	12.5	13.7	9.7	7.6
Net Cash flow after tax	$M	630.0			175.5	170.8	170.6	152.2	131.6	60.9	37.3	30.4	27.6	29.8	13.3	32.3

Disc. cash flow (7.5%)	$M	344.6	(114.2)	(268.2)	151.8	137.5	127.7	106.0	85.3	36.7	21.0	15.8	13.4	13.5	5.6	12.6
Cumulative disc. cash flow	$M		(114.2)	(382.4)	(230.5)	(93.0)	34.7	140.7	226.0	262.7	283.7	299.5	312.9	326.4	332.0	344.6

Net Present Value (7.5%)	$M	344.6
Internal Rate of Return	%	29.5

Total Cash Cost (Gold Equiv.)	US$/oz	700.20
All-in Sustaining Cost (Gold Eq.)	US$/oz	841.57
All-in Cost (Gold Equiv.)	US$/oz	1,046.01

Figure 22.4 Life-of-Mine Base Case Cash Flows

Pre-tax base case cash flows provide an internal rate of return (IRR) of 39.6%; when discounted at the rate of 7.5% per year, the pre-tax net present value (NPV7.5) is $577.7 million.

After-tax cash flows provide an IRR of 29.5%; after-tax NPV7.5 is $344.6 million. Profitability index (i.e., the ratio of NPV7.5/Initial Capital) is 0.9. Undiscounted, the after-tax payback period is 2.3 years. When discounted at 7.5% per year, it extends to 2.7 years.

22.7 SENSITIVITY STUDY AND RISK ANALYSIS

22.7.1 Metal Price and Exchange Rate Assumptions

Micon tested the sensitivity of the base case after-tax NPV7.5 to changes in metal price, operating costs and capital investment for a range of 30% above and below base case values. The impact on NPV7.5 to changes in other revenue drivers such as the grade of material treated and the percentage recovery of metals from processing is equivalent to price changes of the same magnitude, so these factors can be considered as equivalent to the price sensitivity.

Figure 22.5 shows the results of changes in each factor separately. The chart demonstrates that the project remains viable across the range of sensitivity tested. Nevertheless, it is most sensitive to metal price with a reduction of 25% reducing NPV7.5 to $11 million. The project is less sensitive to both operating and capital costs, with an increase of 25% reducing NPV7.5 to $206 million and $208 million, respectively.

Figure 22.5 Sensitivity of Base Case NPV7.5 to Capital, Operating Costs and Metal Prices

Separately, Micon also tested the sensitivity of the Project for specific gold prices above and below the base case price of $1,561/oz. Table 22.4 shows the results of this exercise, demonstrates that each $100/oz change in the gold price results in a change of around $70 million in NPV7.5.

Gold Price	NPV5	NPV7.5	NPV10	IRR	Payback
(US$/oz)	(CDN$M)	(CDN$M)	(CDN$M)	(%)	Disc 7.5% (Yrs)
1,200	171	121	79	16.1%	4.1
1,300	239	181	132	19.9%	3.6
1,400	307	242	187	23.6%	3.2
1,500	379	306	244	27.3%	2.9
1,561	423	345	279	29.5%	2.7
1,700	523	433	358	34.4%	2.4
1,800	591	494	412	37.6%	2.2
1,900	658	553	465	40.6%	2.1
2,000	726	613	518	43.4%	1.9

Table 22.4 Base Case: Sensitivity of NPV, IRR and Payback to Metal Price

Micon notes that in August, 2020 the spot gold price reached a high of more than $2,050/oz, and that the monthly average price has remained above $1,800/oz since then.

22.7.2 Alternative Development Case (Toll Treatment)

Micon evaluated an alternative development option to the base case in which auriferous concentrates are sold for treatment elsewhere. The LOM project cash flow for this tolltreatment scenario is presented in Table 22.5.

	LOM Total$'000	$/t Milled	US$/ozAuEq
Net Revenue	2,044,187	203.61	1,561.00
Mining costs	626,692	62.42	478.56
Processing costs	249,758	24.88	190.72
General & Administrative costs	76,202	7.59	58.19
Total Cash Cost	952,652	94.89	727.47
Net cash operating margin	1,091,536	108.72	833.53
Initial capital	226,631	22.57	173.06
Sustaining capital	273,820	27.27	209.10
Closure bond	6,500	0.65	4.96
Net Cash flow before tax	584,584	58.23	446.41
Taxation	216,960	21.61	165.68
Net Cash flow after tax	367,625	36.62	280.73
All-in Sustaining Cost per ounce (AISC)			936.57
All-in Cost per ounce (AIC)			1,109.63

Table 22.5 Toll Treatment: Life-of-Mine Cash Flow Summary

Pre-tax cash flows in the Toll Treatment scenario provide an IRR of 35.2%; when discounted at the rate of 7.5% per year, the pre-tax net present value (NPV7.5) is $325.9 million.

After-tax cash flows in this scenario provide an IRR of 25.1%; after-tax, compared to 29.5% in the base case. After-tax NPV7.5 is $184.0 million, compared to $344.6 million for the base case. The Profitability index (i.e., the ratio of NPV7.5/Initial Capital) is 0.8 for toll treatment, compared to 0.9 in the base case. Undiscounted, the after-tax payback period is 2.8 years. When discounted at 5% per year, payback extends to 3.3 years.

Thus, by each of these measures, the base case is superior to the alternative.

Annual cash flows for the alternative case are summarized in Figure 22.6.

Figure 22.6 Life-of-Mine Cash Flows (Alternative Case)

Micon tested the sensitivity of the after-tax NPV7.5 to changes in metal price, operating costs and capital investment for a range of 30% above and below base case values. Figure 22.7 shows the results of changes in each factor separately.

Figure 22.7 Sensitivity of NPV7.5 in Alternative Case

The chart demonstrates that this scenario remains viable across the range of sensitivity tested, except that a reduction of 25% in metal prices is seen to reduce NPV7.5 below zero. The project is less sensitive to both operating and capital costs. An increase of 25% to operating costs

reduces NPV7.5 to around $92 million and an increase of 25% to capital costs reduces NPV7.5 to $87 million.

22.8 CONCLUSION

Micon concludes that, based on the forecast production, offtake, capital expenditure and operating cost estimates presented in this study, the project base case demonstrates robust economics across the range of sensitivity tested, with an after-tax NPV7.5 of $344.6 million, IRR of 29.5%, and an all-in sustaining cost (AISC) of US$842/oz gold equivalent. While the alternative case with toll treatment of the auriferous concentrates presents a positive result on base case assumptions, it is economically inferior and less robust when compared to the base case.

23.0 ADJACENT PROPERTIES

The following section is based on the Technical Report on the Property by Puritch et al. (2020).

The Property is situated in a well mineralized area of British Columbia. It is surrounded by several different types of mineralized showings, all within 10 kilometres of the Main Zone portals.

The Mastodon Property is five kilometres to the southeast of the Main Zone. The Mastodon is a group of deposits and showings which include the Mastodon (082M 005), Mastodon North (082M 195), Lead King (082M 094), Little Slide (082M 006) and Little Slide No. 3 (082M 196). The area is a series of polymetallic (Zn, Pb, Cd, Ag, Au, Cu) breccia, replacement-type bodies that are tabular (Mastodon - 90 x 60 x 3 metres) in Badshot Limestone which may be structurally controlled. It displays many of the same characteristics as the Main Zone and could be a parallel mineralized structure. Teck Resources Ltd. ("Teck"), which owns the property, failed to discover sufficient surface indications of mineralization. The entire Mastodon group has had several geochemical surveys completed, with several lead/zinc anomalies having been outlined to-date. Surface drilling of these anomalies has been discouraging.

The Copper Queen showing (082M 004) is 7 kilometres to the southwest of Property. This polymetallic (Cu, Zn, Ag) showing is considered a Kuroko massive sulphide-type deposit (G06). Little work has been done on this deposit to define its overall dimensions.

The Locojo showing (082M 264) is 5 kilometres to the east of the Main Zone. It is a new discovery that has recently been exposed from beneath a glacier. Weymin Mining Corporation was the original group to stake this showing. The showing is considered a Besshi-type massive sulphide (Cu-Zn-Pb) deposit (G04). Very little exploration has been carried out at this showing due to its remote location. The current owner is Imperial Metals Corp.

The authors of this Technical Report have been unable to verify the above information and the mineralization described may not necessarily be representative of the Revel Ridge Property.

24.0 OTHER RELEVANT DATA AND INFORMATION

Neither Micon nor the Authors are aware of any other relevant material data and information that would result in this Technical Report containing misleading statements.

25.0 INTERPRETATION AND CONCLUSIONS

The Property represents one of the largest undeveloped Mineral Resources in British Columbia.

The Property has two known and significant polymetallic precious and base metal deposits. The Main Zone is a structurally controlled orogenic, polymetallic, precious metals enhanced gold deposit. The sheeted massive sulphide Main Zone system is composed of banded massive and stringer arsenopyrite-pyrite-sphalerite-galena vein like tabular mineralization with appreciable content of gold and silver. The Main Zone has been traced on surface for a strike length of over three kilometres and traced by drilling for 1,500-metres of strike length and 800 metres down dip. The Main Zone generally dips about 56° to the northeast with an average true thickness of 2.5 metres but can reach 15 metres true thickness.

The silver-lead-zinc-rich Yellowjacket Zone is considered to be a structurally controlled carbonate replacement deposit composed of multiple parallel siliceous sphalerite-galenabearing zones. The individual zones making up the Yellowjacket Zone occur as lenticular bodies each up to eight metres thick at the contact between alternating units of phyllites and limestone. The Yellowjacket Zone sub parallels and is in the immediate hanging wall of the Main Zone. The Yellowjacket Zone has no notable gold but has higher silver, lead and zinc values than the Main Zone. It is open along strike and down dip. The very high silver contents of the Yellowjacket zone are highly anomalous with respect to other Kootenay Arc Pb-Zn deposits.

The Property has been explored by a number of mining companies by trenching, tunnelling and drilling. There is a total of at least 315 drill holes that have been completed on the Property from 1983 to present. This translates to 41,075.9 metres of drilling. The 830 m level drift and related crosscuts total 3.1 kilometres exposing the Main Zone for approximately 800 metres. The 550-metre long 832 trackless drift provides year-round underground access to the 830 drift.

Underground bulk samples have been taken from the Main Zone to conduct metallurgical testwork. The Main Zone is a complex polymetallic deposit high in arsenic values which create a challenge in the production of saleable zinc and lead concentrates and the economic recovery of gold. Extensive metallurgical testing between the mid 1980s and 2014 have considered various options and have produced numerous effective options for acceptable recoveries of gold, silver, zinc and lead by making three separate concentrates. Based on the current envisioned circuit and corresponding laboratory test response, the overall process recoveries for the Main Zone are expected to be approximately 93% Au, 70% Ag, 74% Pb, and 80% Zn. Limited metallurgical testwork from drill core has been performed on the Yellowjacket Zone which has less complex metallurgy than the Main Zone. The expected process recoveries for the Yellowjacket Zone are 94% Ag, 88% Pb, and 93% Zn.

The Qualified Person is satisfied that the drill sample database and geological interpretations are sufficient to enable the estimation of Mineral Resources. Accepted estimation methods

have been used in the generation of a 3-D block model of Au, Ag, Pb and Zn grades and assigned bulk densities.

The Mineral Resource Estimates have been classified with respect to CIM Standards as Measured, Indicated and Inferred, according to the geological confidence and sample spacings that currently define the deposit. In the case of the Main Zone, Measured Mineral Resources require 30 metre drill centres. Indicated Mineral Resources require 60 metre drill centres and Inferred Mineral Resources require 120 metre drill centres.

The Author is of the opinion that the current Mineral Resource Estimate has a reasonable prospect of economic extraction due to its approximate 7.0 g/t AuEq average grade and the $110/t NSR cut-off (equal to approx. 3.5 g/t AuEq). The Author has experience with other similar projects and is of the opinion that the cut-off grade and cost assumptions are reasonable.

The Author is not aware of any environmental, permitting, legal, title, taxation, socioeconomic, marketing, political, or other relevant factors which may materially affect the Mineral Resource estimate. A material drop in metal prices below the Dec 31, 2019 two year trailing average prices used for the current Mineral Resource Estimate or a significant increase in operating costs could materially affect the cut-off and average grades and potentially result in a revised lower Mineral Resource Estimate tonnage.

This Preliminary Economic Assessment is preliminary in nature; it includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be classified as Mineral Reserves, and there is no certainty that the Preliminary Economic Assessment will be realized.

Pre-tax base case cash flows provide an IRR of 39.6%; when discounted at the rate of 7.5% per year, the pre-tax net present value (NPV7.5) is $577.7 million. After-tax cash flows provide an IRR of 29.5%; after-tax NPV7.5 is $344.6 million. Profitability index (i.e., the ratio of NPV7.5/Initial Capital) is 0.9. Undiscounted, the after-tax payback period is 2.3 years. When discounted at 7.5% per year, it extends to 2.7 years.

The project is most sensitive to gold price with a reduction of 25% reducing NPV5 to $60 million. The project is less sensitive to both operating and capital costs, with an increase of 25% reducing NPV5 to $166 million and $172 million, respectively.

26.0 RECOMMENDATIONS

The authors make the following recommendations for future work:

There is potential to expand both the Main Zone and Yellowjacket beyond their currently dimensions as defined by drilling. The Main Zone, in particular, with its tabular predictable geometry and grade, has already a laterally extensive size defined by drilling and remains open in a number of directions. The down dip and strike towards the southeast on the Main Zone hold the best potential to build additional Mineral Resources.
A program to advance the Revel Ridge Project through a Pre-Feasibility Study would be appropriate at an estimated cost of $800,000. Associated with the Pre-Feasibility Study additional recommended work includes metallurgy, geotechnical site assessment drilling, First Nations consultation and environmental studies. These additional studies are estimated to cost an additional $1,400,000. Additionally, a 16,000-metre diamond drill program should be conducted as part of an ongoing Mineral Resource expansion program at a cost of $4,800,000.
• The proposed budget for the PFS program is presented in Table 26.1.

Task Description	Cost (CDN$)
Preliminary Feasibility Study
Metallurgical Testwork	750,000
Geotechnical Mine & Site Assessment Drilling	400,000
Environmental Study Initiation	250,000
Diamond drilling (16,000 m)	4,800,000
Pre-Feasibility Study	800,000
PFSSubtotal	7,000,000
Contingency at 15%	1,050,000
PFSTotal	8,050,000

Table 26.1 Budget for Proposed PFS Program

27.0 DATE AND SIGNATURE PAGE

"Eugene Puritch" {signed, and sealed}

Eugene Puritch, P.Eng., FEC, CET January 22, 2021

"Fred Brown" {signed and sealed}

Fred Brown, P.Geo. January 22, 2021

"Jarita Barry" {signed and sealed}

Jarita Barry, P.Geo. January 22, 2021

"Richard Routledge" {signed and sealed}

Richard Routledge, P.Geo. January 22, 2021

"Nigel S. Fung" {signed and sealed}

Nigel S. Fung, P.Eng. January 22, 2021

"Richard M. Gowans" {signed and sealed}

Richard M. Gowans, P.Eng. January 22, 2021

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Makepeace, D.K., 2007. J&L Property Technical Report (43-101) for Merit Mining Corp.

McClay, K.R., 1984. The structure of the J&L Polymetallic Sulphide Deposit, British Columbia, Private Report for BP Canada Ltd., Selco Division.

McKinlay, F.T., 1987. Geology and Control of Sulphide Deposition of the J&L Massive Sulphide Deposit, Southeast British Columbia; unpublished M.Sc. thesis, The University of Western Ontario.

Muraro, T.W., 1966. Metamorphism of Zinc-Lead Deposits in Southeastern British Columbia in Tectonic History and Mineral Deposits of the Western Cordillera: Canadian Institute of Mining and Metallurgy, Special Volume 8, pp. 239-247.

Meyers, R.E., Hubner T.B., 1989. An Update on the J&L Gold-bearing Polymetallic Sulphide Deposit; in Exploration in British Columbia, BCMEMPR, pp 81-89.

Oliver, J.L., 1990. Geological Evolution of the J&L Gold-Silver-Lead Zinc Property, Revelstoke Mining Division 82M/8E Private Corporate Report for Placer Dome Inc., 49 p.

Paradis, S. and Simandl, G.J. 2010. Carbonate-hosted sulphide and non-sulphide Pb-Zn mineralization, British Columbia, Canada, focus on new Exploration Criteria. TGI-3

Workshop: Public geoscience in support of base metal exploration programme and abstracts; Geological Association of Canada, Cordilleran Section; p. 60-62.

Pegg, R., Jan.1983. A Summary Report on the J&L Option, Lead-Zinc-Gold-Silver Prospect, British Columbia, NTS 82M/8E, Private Corporate Report for BP-Selco Inc., 160 p.

Pegg, R., Grant, B., March 1984. A Summary Report on the J&L Option, Lead-Zinc-Gold-Silver Prospect, British Columbia, NTS 82M/8E, Private Corporate Report for BP-Selco Inc., 72 p.

Pegg, R., Grant, B., Feb.1985. A Summary Report on the J&L Option, Lead-Zinc-Gold-Silver Prospect, British Columbia, NTS 82M/8E, Private Corporate Report for BP-Selco Inc., 66 p.

Pegg, R., Dec.1985. A Summary Report on the J&L Option, Lead-Zinc-Gold-Silver Prospect, British Columbia, NTS 82M/8E, Private Corporate Report for BP-Selco Inc., 55 p.

Pegg, R., (1982-1985). Assessment Reports Related to Various Physical, Geological, Geophysical and Geochemical Surveys Carried Out on the J&L Property for BP-Selco Inc., (Assessment Reports: 10939, 12616, 12634 and 14405).

Puritch, E, Brown, F., Hayden, A., Barry, J., Routledge, R., 2020. Updated Technical Report on the Revel Ridge Property (Formerly J&L Property), Revelstoke Mining Division, British Columbia, for Rokmaster Resources Corp., P&E Mining Consultants Inc.

Puritch, E, Brown, F., Hayden, A., Barry, J., Routledge, R., 2018. Technical Report and Updated Mineral Resource Estimate, on the J&L Property, Revelstoke, British Columbia, for Golden Dawn Minerals Inc., P&E Mining Consultants Inc.

Riddel, J.E., 1946, Preliminary Report on Raindor Gold Mines Limited.

Smith & Dvorak, 1982, Dighem II Survey of the J&L Prospect.

Squair, H., 1981. A Report on the J and L Lead-Zinc-Gold-Silver Prospect for Selco Mining Corp.

Starr, C.C., 1926, Report on Preliminary Examination of the J&L Mine, Revelstoke, BC.

Starr, C.C., 1928, Report of Examination of the J&L Mine, Revelstoke, BC.

Sullivan, J., 1967. Report on Westairs Mines Ltd. J&L Project. A private corporate report for Westairs Mines Limited.

Timmins, W.G., 1979, Geological Report on the J&L Project, Private Report for Stelladord Mines Ltd.

Weicker, R., 1989. A Summary Report on A&E Showings, J&L Property.

Private Corporate Report for Equinox Resources Ltd. (see Assessment Report No. 19454).

Weicker, R., 1990. Geochemistry and Hydrology Report on Carnes and McKinnon Creeks for Equinox Resources Ltd. (Assessment Report No. 20716).

Weicker, R., 1991. Report on 1991 Summer Exploration Program, J&L Property for Equinox Resources Ltd. (see Assessment Report No. 22004).

Weicker,R., 1991. Report on 1990-1991 Exploration Program, J&L Property for Cheni Gold Mines Inc.

Wheeler, J.O., 1964. Geology of the Big Bend Map Area, British Columbia, Geological Survey of Canada, Paper 64-32, 37 p.

Wright, J.H., Weicker, R.F., 1989. Completion Report on Phase I Exploration Program J&L Property, BC. Unpublished Report for Equinox Resources Ltd.

Wright, J.H., Weicker. R., Taal, T., 1989. Diamond Drilling and Metallurgical Testwork on the J&L Property for Equinox Resources Ltd. (Assessment Report No. 19469) 59 p.

Wynne, A., 1982. Summary Report of the Questor and Dighem Airborne Electromagnetic Surveys.

Wynne, A., 1983. J&L Project – A Report on Ground Geophysical Survey.

Metallurgical References (Chronological)

1982

Stairs, I., 1982. Summary Report on the J&L Property for Selco Inc.

Mineralogical Examination of J&L Project – Samples Submitted by Selco Mining Corp. Progress Report No. 1 and No.2 by Lakefield Research of Canada Ltd.

An Investigation of the Recovery of Lead, Zinc and Precious Metals from Ore Samples submitted by Selco Inc., Progress Report No.3 by Lakefield Research.

1983

An Investigation of the Recovery of Lead, Zinc and Precious Metals from Ore Samples submitted by Selco Inc., Progress Report No.4 by Lakefield Research.

1984

An Investigation of the Recovery of Lead, Zinc and Precious Metals from Ore Samples submitted by Selco Inc., Progress Report No.5 by Lakefield Research.

1985

Bennett, C.A., Preliminary Metallurgical Evaluation of Six Ore Types from the J&L Property, (Technical Memorandum 134 731), for BP Selco.

Kim, Moon, Review of the Metallurgical Investigation on J&L Ores.

Lichty, L.J., Preliminary Report of Metallurgical Testwork on Lead-Zinc-Gold- Silver Ore from the J&L Property for Pan American Minerals Corp.

Mineralogical Examination of Lead-Zinc Samples – Submitted by BP Canada (Selco) Ltd. Progress Report No.6 by Lakefield Research.

Schedler, R.A., Johnston, D.C., J&L Property: Mineralogical Study of Crushed Samples of the Six Main Ore Types.

1987

Stowe, K., Metallurgical Assistance – J&L – Report No.1 Centre de Recherche Noranda.

Lichty, L.J., Interim Report No.1, The Treatment of J&L Ore of Pan American Minerals Corp. by the Cashman Process (Mountain States Laboratory, Tuscon, Arizona).

Preliminary Bio-oxidation Tests on J&L Arsenopyrite Concentrates by R.C. Smith and Associates.

The Metallurgical Response of Pan American J&L Ore, Met Engineers Ltd.

Preliminary Metallurgical Testing of a Sample from J&L Polymetallic Deposit by Bacon, Donaldson and Associates for Pan American Minerals Corp.

1988

Lichty, L.J., Interim Report No.2, The Treatment of J&L Ore of Pan American Minerals Corp. by the Cashman Process.

J&L Property – Metallurgy by R.C. Smith and Associates for Pan American Minerals Corp.

An Investigation of Samples of Bulk Concentrates submitted by Pan American Minerals Corp., Artech Recovery Systems Inc.

1989

Pilot Plant Program - J&L Property, Progress Report No.1 for Equinox Operations Group by Lakefield Research.

An Investigation of the Recovery of Gold from Flotation Products submitted by Equinox Resources Ltd., Progress Report No. 2 by Lakefield Research.

An Investigation of the Recovery of Lead-Zinc-Gold and Silver from J&L Drill Hole Composite Samples for Equinox Resources Ltd., Progress Report No.1 by Lakefield Research.

Preliminary Study Order of Magnitude Capital and Operating Cost estimate Pressure Oxidation Treatment of J&L Refractory Material for Equinox Resources Ltd., by Wright Engineers Limited.

1990

An Investigation of the Recovery of Lead-Zinc-Gold and Silver from J&L Composite Samples for Equinox Resources Ltd., Progress Report No.1 by Lakefield Research.

An Investigation of the Recovery of Lead-Zinc-Gold and Silver from J&L Composite Samples for Equinox Resources Ltd., Progress Report No.3 by Lakefield Research.

Hottin, G., Evaluation du Potentiel du Gite de J&L by BRGM.

Vincent, J.S., An Executive Summary on the J&L Polymetallic Gold Project.

A Pilot Plant Investigation of the Recovery of Lead, Zinc, Gold and Silver from J&L Composite Samples submitted by Equinox Resources Ltd., by Lakefield Research.

1991

Zinc Recovery from a Redox Process Bleed Stream by Solvent or Ion Exchange Extraction for Cheni Gold Mines Ltd., by Bacon Donaldson and Associates.

An Investigation of the Separation of Arsenopyrite and Pyrite from J&L Project Samples, submitted by Cheni Gold Mines Ltd., Progress Report No.1, by Lakefield Research.

Results of Redox Process Testing of J&L Property Flotation Concentrates for Cheni Gold Mines by Bacon Donaldson and Associates.

Bio-oxidation of J&L Refractory Gold-Bearing Sulphide Concentrate for Cheni Gold Mines by Coastech Research Inc.

Gochin, R.J., Diaz, M.A., The Separation of Pyrite and Arsenopyrite by Flotation – Progress Report No.1, Department of Mineral Resources Engineering, Imperial College of Science, Technology and Medicine, London, England.

Durston, K, Review and Evaluation of the J&L Property by Cominco Engineering Services Ltd. (CESL).

The Flotation Concentration of J&L Complex Sulphide Ore for Cheni Gold Mines Ltd., by Bacon Donaldson and Associates.

Variability Testing of Samples from J&L Deposit for Cheni Gold Mines Ltd, by Bacon Donaldson and Associates.

Heavy Medium Separation for J&L Yellowjacket Samples for Cheni Gold Mines Ltd., by Bacon Donaldson and Associates.

1996

Project Opportunities for J&L Property for Weymin Resources Ltd. by H. A.Simons.

Technical Review of J&L Property for Weymin Resources Ltd. by H.A. Simons.

1998

J&L project – Heavy Media Separation Study for Weymin Mining Corp. by Process Research Associates Ltd.

Preliminary Flotation Testing of Yellowjacket Drill Core for Weymin Mining Corp. by Beattie Consulting Ltd.

Preliminary Report – Weymin Mining Corp. - McKinnon Creek Project, Hydrometallurgical Scoping by March Process Consulting Ltd.

J&L Project – Metallurgical Test Report for Weymin Mining Corp by Process Research Associates Ltd.

J&L Project Update – Progress on Metallurgical Testwork by Beattie Consulting Ltd.

McKinnon Creek Property – Scoping Study for Weymin Mining Corp. by H.A. Simons Ltd.

2005

McKinnon Creek Project – Metallurgical Test Program Results, for BACTECH Mining Corporation by Process Research Associates Ltd.

2007

J&L Property, Preliminary Project Evaluation for Merit Mining Corp. by Dynatec Corp.

Gormley, L. (AMEC), Fatal Flaw Review of Metallurgical Testwork and Data for Merit Mining Corp.

2011

Lang, J., Review of Historical Metallurgical Results for J&L Project by SGS Canada Inc.

2014

Wright, F., J&L Project Metallurgical Response for Huakan International Mining Inc. by F. Wright Consulting. Includes appendices detailing metallurgical testwork undertaken between 2011 and 2013 at Inspectorate Exploration and Mining Services Ltd. of Richmond, BC, Hazen Research in Golden CO., and SGS in Lakefield ON.

2020

Metallurgical Testing of the Revel Ridge Gold Project – BL0604 by Base Metallurgical Laboratories.

Environmental References

Equinox Operations Group, 1990. Geochemistry and Hydrology Report on Carnes and McKinnon Creeks. Revelstoke Mining Division. British Colombia.

Haukan International Mining, 2020. Annual Reclamation Report for 2019 on the J&L Property.

BC Ministry of Environment & Climate Change Strategy, 2017. Technical Guidance on Contaminated Sites: Concentration Limits for the Protection of Aquatic Receiving Environments. https://www2.gov.bc.ca/assets/gov/environment/air-land-water/siteremediation/docs/technical-guidance/tg15.pdf?bcgovtm=CSMLS

BC Ministry of Environment & Climate Change Strategy, 2019. British Columbian Approved Water Quality Guidelines: Aquatic Life, Wildlife & Agriculture https://www2.gov.bc.ca/assets/gov/environment/air-land-water/water/waterquality/waterquality-guidelines/approved-wqgs/wqg\_summary\_aquaticlife\_wildlife\_agri.pdf

29.0 CERTIFICATES

CERTIFICATE OF QUALIFIED PERSON EUGENE PURITCH, P. ENG., FEC, CET

I, Eugene J. Puritch, P. Eng., FEC, CET, residing at 44 Turtlecreek Blvd., Brampton, Ontario, L6W 3X7, do hereby certify that:

1. I am an independent mining consultant and President of P&E Mining Consultants Inc.
1. This certificate applies to the Technical Report titled "An Updated Preliminary Economic Assessment of the Revel Ridge Project, Revelstoke, B.C., Canada" (The "Technical Report") with an effective date of December 08, 2020.
1. I am a graduate of The Haileybury School of Mines, with a Technologist Diploma in Mining, as well as obtaining an additional year of undergraduate education in Mine Engineering at Queen's University. In addition, I have also met the Professional Engineers of Ontario Academic Requirement Committee's Examination requirement for Bachelor's Degree in Engineering Equivalency. I am a mining consultant currently licensed by the: Professional Engineers and Geoscientists New Brunswick (License No. 4778); Professional Engineers, Geoscientists Newfoundland and Labrador (License No. 5998); Association of Professional Engineers and Geoscientists Saskatchewan (License No. 16216); Ontario Association of Certified Engineering Technicians and Technologists (License No. 45252); Professional Engineers of Ontario (License No. 100014010); Association of Professional Engineers and Geoscientists of British Columbia (License No. 42912); and Northwest Territories and Nunavut Association of Professional Engineers and Geoscientists (No. L3877). I am also a member of the National Canadian Institute of Mining and Metallurgy.

I have read the definition of "Qualified Person" set out in National Instrument 43-101 ("NI 43-101") and certify that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "Qualified Person" for the purposes of NI 43-101.

I have practiced my profession continuously since 1978. My summarized career experience is as follows:

Mining Technologist -H.B.M.& S. and Inco Ltd.,	1978-1980
Open Pit Mine Engineer –Cassiar Asbestos/Brinco Ltd.,	1981-1983
Pit Engineer/Drill & Blast Supervisor –Detour Lake Mine,	1984-1986
Self-Employed Mining Consultant –Timmins Area,	1987-1988
Mine Designer/Resource Estimator –Dynatec/CMD/Bharti,	1989-1995
Self-Employed Mining Consultant/Resource-Reserve Estimator,	1995-2004
President –P&E Mining Consultants Inc,	2004-Present

1. I have not visited the Property that is the subject of this Technical Report.
1. I am responsible for authoring Sections 2, 3 and co-authoring Sections 1, 14, 24, 25 and 26 of this Technical Report.
1. I am independent of the Issuer applying the test in Section 1.5 of NI 43-101. I am independent of the Vendor and the Property.
1. I have had prior involvement with the Project that is the subject of this Technical Report. Most recently, I was a "Qualified Person" for a Technical Report titled "Updated Technical Report on The Revel Ridge Property (Formerly J&L Property)", (The "Technical Report") with an effective date of January 29, 2020.
1. I have read NI 43-101 and Form 43-101F1. This Technical Report has been prepared in compliance therewith.
1. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: December 08, 2020 Signed Date: January 22, 2021

"Eugene Puritch" {SIGNED AND SEALED}

____________________________ Eugene Puritch, P.Eng., FEC, CET

CERTIFICATE OF QUALIFIED PERSON FRED H. BROWN, P.GEO.

I, Fred H. Brown, of PO Box 332, Lynden, WA, USA, do hereby certify that:

1. I am an independent geological consultant and have worked as a geologist continuously since my graduation from university in 1987.
1. This certificate applies to the Technical Report titled "An Updated Preliminary Economic Assessment of the Revel Ridge Project, Revelstoke, B.C., Canada" (The "Technical Report") with an effective date of December 08, 2020.
1. I graduated with a Bachelor of Science degree in Geology from New Mexico State University in 1987. I obtained a Graduate Diploma in Engineering (Mining) in 1997 from the University of the Witwatersrand and a Master of Science in Engineering (Civil) from the University of the Witwatersrand in 2005. I am registered with the South African Council for Natural Scientific Professions as a Professional Geological Scientist (registration number 400008/04), the Association of Professional Engineers and Geoscientists of British Columbia as a Professional Geoscientist (171602) and the Society for Mining, Metallurgy and Exploration as a Registered Member (#4152172).

My relevant experience for the purpose of the Technical Report is:

Underground Mine Geologist, Freegold Mine, AAC	1987-1995
Mineral Resource Manager, Vaal Reefs Mine, Anglogold	1995-1997
Resident Geologist, Venetia Mine, De Beers	1997-2000
Chief Geologist, De Beers Consolidated Mines	2000-2004
Consulting Geologist	2004-2008
P&E Mining Consultants Inc. –Sr. Associate Geologist	2008-Present

1. I have visited the Property that is the subject of this Technical Report on December 17, 2010.
1. I am responsible for co-authoring Sections 1, 12, 14, 25 and 26 of this Technical Report.
1. I am independent of the Issuer applying the test in Section 1.5 of NI 43-101. I am independent of the Vendor and the Property.
1. I have had prior involvement with the Project that is the subject of this Technical Report. Most recently I was a "Qualified Person" for a Technical Report titled "Updated Technical Report on The Revel Ridge Property (Formerly J&L Property)", (The "Technical Report") with an effective date of January 29, 2020.
1. I have read NI 43-101 and Form 43-101F1 and this Technical Report has been prepared in compliance therewith.
1. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: December 08, 2020 Signed Date: January 22, 2021

____________________________

"Fred H. Brown" {SIGNED AND SEALED}

Fred H. Brown, P.Geo.

CERTIFICATE OF QUALIFIED PERSON JARITA BARRY, P.GEO.

I, Jarita Barry, P.Geo., residing at 4 Creek View Close, Mount Clear, Victoria, Australia, 3350, do hereby certify that:

1. I am an independent geological consultant contracted by P&E Mining Consultants Inc.
1. This certificate applies to the Technical Report titled "An Updated Preliminary Economic Assessment of the Revel Ridge Project, Revelstoke, B.C., Canada" (The "Technical Report") with an effective date of December 08, 2020.
1. I am a graduate of RMIT University of Melbourne, Victoria, Australia, with a B.Sc. in Applied Geology. I have worked as a geologist for a total of 13 years since obtaining my B.Sc. degree. I am a geological consultant currently licensed by Engineers and Geoscientists British Columbia (License No. 40875), Professional Engineers and Geoscientists Newfoundland & Labrador (License No. 08399) and Northwest Territories and Nunavut Association of Professional Engineers and Geoscientists (License No. L3874). I am also a member of the Australasian Institute of Mining and Metallurgy of Australia (Member No. 305397).

I have read the definition of "Qualified Person" set out in National Instrument 43-101 ("NI 43-101") and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "Qualified Person" for the purposes of NI 43-101.

My relevant experience for the purpose of the Technical Report is:

Geologist, Foran Mining Corp.	2004
Geologist, Aurelian Resources Inc.	2004
Geologist, Linear Gold Corp.	2005-2006
Geologist, Búscore Consulting	2006-2007
Consulting Geologist (AusIMM)	2008-2014
Consulting Geologist, P.Geo. (APEGBC/AusIMM)	2014-Present

1. I have not visited the Property that is the subject of this Technical Report.
1. I am responsible for authoring Sections 4 to 11, 23 and co-authoring Sections 1, 12, 25 and 26 of this Technical Report.
1. I am independent of the Issuer applying the test in Section 1.5 of NI 43-101. I am independent of the Vendor and the Property.
1. I have had prior involvement with the Project that is the subject of this Technical Report. Most recently I was a "Qualified Person" for a Technical Report titled "Updated Technical Report on The Revel Ridge Property (Formerly J&L Property)", (The "Technical Report") with an effective date of January 29, 2020.
1. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith.
1. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: December 08, 2020 Signed Date: January 22, 2021

"Jarita Barry" {SIGNED AND SEALED}

________________________________

Jarita Barry, P.Geo.

CERTIFICATE OF QUALIFIED PERSON RICHARD E. ROUTLEDGE, P.GEO.

I, Richard E. Routledge, P.Geo., residing at 1386 Queen's Line, PO Box 335, Minden, Ontario, K0M 2K0, do hereby certify that:

1. I am an independent Consulting Geologist who has been contracted by P&E Mining Consultants Inc.
1. This certificate applies to the Technical Report titled "An Updated Preliminary Economic Assessment of the Revel Ridge Project, Revelstoke, B.C., Canada" (The "Technical Report") with an effective date of December 08, 2020.
1. I graduated with a Bachelor of Science degree, Major in Geology, from Sir George Williams (Concordia) University in 1971 and with a Masters degree in Applied Exploration Geology from McGill University in 1973. I have worked as a geologist for about 43 years since post-graduation. I am a Professional Geologist registered in the Province of Ontario (APGO No. 1354).

I have read the definition of "qualified person" set out in National Instrument 43-101 ("NI 43-101") and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "Qualified Person" for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:

Independent Consulting Geologist.	2011 –Present
Roscoe Postle Associates Inc., Consulting Geologist	1998 –2011
Independent Consulting Geologist	1994 –1997
Vice President Exploration, Greater Lenora Resources Corp.	1993 –1994
Teck Explorations Ltd, Evaluations and Mineral Commodities Geologist.	1985 –1992
Derry, Michener, Booth & Wahl, Exploration and Consulting Geologist.	1973 –1985

1. I have visited the property that is the subject of this Technical Report on June 13 and 14, 2012.
1. I am responsible for co-authoring Sections 1, 12, 25 and 26 of this Technical Report.
1. I am independent of the Issuer applying the test in Section 1.5 of NI 43-101. I am independent of the Vendor and the Property.
1. I have had prior involvement with the Project that is the subject of this Technical Report. Most recently I was a "Qualified Person" for a Technical Report titled "Updated Technical Report on The Revel Ridge Property (Formerly J&L Property)", (The "Technical Report") with an effective date of January 29, 2020.
1. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith.
1. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report misleading.

Effective Date: December 08, 2020 Signed Date: January 22, 2021

________________________________

"Richard E. Routledge" {SIGNED AND SEALED}

Richard E. Routledge, P. Geo.

CERTIFICATE OF QUALIFIED PERSON NIGEL S. FUNG, P.ENG.

As an author of this report titled "An Updated Preliminary Economic Assessment of the Revel Ridge Project, Revelstoke, B.C., Canada" (The "Technical Report") with an effective date of December 08, 2020 (the "Technical Report"), I, Nigel S. Fung do hereby certify that:

1. I am employed by, and carried out this assignment for, Micon International Limited, 900 390 Bay Street, Toronto, Ontario M5H 2Y2, tel. (416) 362-5135, e-mail [email protected].
1. I hold the following academic qualifications:

B.Sc. (H) Biology, The University of Toronto 1993

B.Eng. (Mining), McGill University 2001

1. I am a registered Professional Engineer in Ontario (PEO membership number 100173276); I am also a registered Engineer with the Society for Mining, Metallurgy & Exploration (SME) (#4185435).
1. I am familiar with NI 43-101 and by reason of education, experience and professional registration and fulfill the requirements of a Qualified Person as defined in NI 43-101. I have been employed as a mining engineer in the minerals industry for over 12 years. My work experience includes 12 years in mine planning in oil sands, gold, and base metals, with over 10 years in open-pit and two years in underground mines. While working with Caterpillar, I have carried out mining production studies throughout Africa, Central Asia and the Middle East in Coal, Precious Metal and industrial mineral mines.
1. I have not visited the Property that is the subject of this Technical Report.
1. I am responsible for Sections 15, 16, 21.1.1, 21.2.2 and 22 and summaries therefrom in Sections 1, 25 and 26 of this Technical Report.
1. I have had prior involvement with the Project that is the subject of this Technical Report.
1. As of the date of this Certificate, 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.
1. I am independent of the Issuer, the Vendor and the Property as defined by Section 1.5 of the Instrument.
10.I have read National Instrument 43-101 and the Technical Report has been prepared in compliance with National Instrument 43-101 and Form 43-101F1.

Effective Date: December 08, 2020 Signed Date: January 22, 2021

"Nigel S. Fung" {SIGNED AND SEALED}

________________________________

Nigel S. Fung, P.Eng.

CERTIFICATE OF QUALIFIED PERSON RICHARD M. GOWANS, P.ENG.

1. I am employed by, and carried out this assignment for, Micon International Limited, 900 390 Bay Street, Toronto, Ontario M5H 2Y2, tel. (416) 362-5135, e-mail [email protected].
1. I hold the following academic qualifications:

B.Sc. (Hons) Minerals Engineering, The University of Birmingham, U.K. 1980

1. I am a registered Professional Engineer in Ontario (PEO membership number 90529389); as well, I am a member in good standing of the Canadian Institute of Mining, Metallurgy and Petroleum.
1. I am familiar with NI 43-101 and by reason of education, experience and professional registration and fulfill the requirements of a Qualified Person as defined in NI 43-101. I have been continuously employed in the mining industry since graduation and my work experience includes over 30 years of the management of technical studies and design of numerous metallurgical testwork programs and metallurgical processing plants.
1. I have not visited the Property that is the subject of this Technical Report.
1. I am responsible for Sections 13, 17, 18, 19, 20, 21 (except for 21.1.1 and 21.2.2) and summaries therefrom in Sections 1, 25 and 26 of this Technical Report.
1. I have had no prior involvement with the Project that is the subject of this Technical Report.
1. As of the date of this Certificate, 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.
1. I am independent of the Issuer, the Vendor and the Property as defined by Section 1.5 of the Instrument.
10.I have read National Instrument 43-101 and the Technical Report has been prepared in compliance with National Instrument 43-101 and Form 43-101F1.

Effective Date: December 08, 2020 Signed Date: January 22, 2021

________________________________

"Richard M. Gowans" {SIGNED AND SEALED}

Richard M. Gowans P.Eng.

StopeLevel	Y3	Y4	Y5	Y6	Y7	Y8	Y9	Y10	Y11	Y12
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660	R	M					M			M
640	R	M			M	M
620		R	M			M
600		R	M			M
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StopeLevel	Y3	Y4	Y5	Y6	Y7	Y8	Y9	Y10	Y11	Y12
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660	R	M					M			M
640	R	M			M	M
620		R	M			M
600		R	M			M
580		R	M				M
560			R	M			M
540				R				M	M
520								R	M
500								R	M

AI assistant

Rokmaster Resources Corp. — Regulatory Filings 2021

46864_rns_2021-01-22_634145e7-38d2-4db5-a41b-5ce07a624e1a.pdf

ROKMASTER RESOURCES CORP.

NI 43-101 TECHNICAL REPORT

AN UPDATED PRELIMINARY ECONOMIC ASSESSMENT OF THE

REVEL RIDGE PROJECT, REVELSTOKE, B.C., CANADA

List of Tables

Page

List of Figures

Page

1.0 SUMMARY

2.0 INTRODUCTION AND TERMS OF REFERENCE

2.1 SOURCES OF INFORMATION

2.2 UNITS AND CURRENCY

2.3 GLOSSARY AND ABBREVIATION OF TERMS

Table 2.1 List of Abbreviations

3.0 RELIANCE ON OTHER EXPERTS

4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 PROPERTY LOCATION

4.2 HUAKAN – ROKMASTER AGREEMENT TERMS

4.3 PROPERTY DESCRIPTION

4.4 GENERAL REQUIREMENTS FOR MINERAL CLAIMS

4.5 PERMITTING

4.6 ARMEX STATEMENT OF CLAIM

4.7 FIRST NATIONS WITH POTENTIAL INTERESTS IN THE REVELSTOKE REGION

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 ACCESSIBILITY

5.2 CLIMATE

5.3 LOCAL RESOURCES

5.4 INFRASTRUCTURE

5.5 PHYSIOGRAPHY

6.0 HISTORY

6.1 PREVIOUS MINERAL RESOURCE ESTIMATES

Table 6.4 Previous 2012 Mineral Resource Estimate

7.0 GEOLOGICAL SETTING AND MINERALIZATION

7.1 REGIONAL GEOLOGY

7.2 PROPERTY GEOLOGY

7.2.1 Hamill Group

7.2.2 Mohican Formation

Figure 7.2 Revel Ridge Regional Geology Map

7.2.3 Badshot Formation

7.3 LARDEAU GROUP

7.3.1 Index Formation

7.3.2 Micaceous Quartzite Unit

7.3.3 Jowett Formation

7.4 LOCAL GEOLOGY

7.5 STRUCTURE

7.6 MINERALIZATION

7.6.1 Main Zone

7.6.2 Hanging Wall and Footwall Zones

7.6.3 Yellowjacket Zone

7.6.4 Other Showings

8.0 DEPOSIT TYPES

8.1 MAIN ZONE

8.2 YELLOWJACKET ZONE

9.0 EXPLORATION

10.0 DRILLING

Table 10.1 Drill Programs Summary

10.1 HUAKAN DIAMOND DRILLING

10.1.1 Collar Surveying

10.1.2 Downhole Surveying

10.1.3 Core Recovery and Storage

10.1.4 Drill Core Size and Orientation

10.1.5 Contractor

10.1.6 Comments

11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

11.1 SAMPLING METHOD AND APPROACH BY HUAKAN

11.2 CHAIN OF CUSTODY

11.3 SAMPLE PREPARATION AND ANALYSES

11.4 HUAKAN QUALITY ASSURANCE/QUALITY CONTROL

11.4.1 Certified Reference Material

11.4.1.1 Blanks

11.4.1.2 Duplicate Sampling Program

12.0 DATA VERIFICATION

12.1 2010/2011 SITE VISIT AND INDEPENDENT SAMPLING

12.2 2012 SITE VISIT AND INDEPENDENT SAMPLING

12.3 DATABASE VERIFICATION

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 INTRODUCTION

Rokmaster Resources Corp. Regulatory Filings 2021

StopeLevel	Y3	Y4	Y5	Y6	Y7	Y8	Y9	Y10	Y11	Y12
680	R	M					M
660	R	M					M			M
640	R	M			M	M
620		R	M			M
600		R	M			M
580		R	M				M
560			R	M			M
540				R				M	M
520								R	M
500								R	M