ASX Announcement

8 December 2025

Gonneville Palladium-Nickel-Copper Project PFS

Pre-Feasibility Study confirms a long life, globally competitive critical minerals mine in Western Australia, set to generate A$4.7bn in free cashflow pre-tax over an initial 23 year open-pit mine life, with a rapid payback of 2.7 years

Highlights

« Pre-Feasibility Study (PFS) completed and maiden Ore Reserve defined for the 23-year open-pit phase of the 100%-owned Gonneville Palladium-Nickel-Copper Project, ~70km from Perth in WA.
« Initial 23 year, two-stage open-pit phase has robust unleveraged financial metrics using base case price assumptions of Pd: US$1,300/oz, Ni: US$18,750/t and Cu: US$10,500/t[1] :
« Cumulative free cashflow pre-tax of A$4.7bn , increasing to A$6.2bn at spot prices[2] ,
« Pre-tax NPV8% of A$1.4bn , increasing to A$2.0bn at spot,
« Post-tax NPV8% of A$1.0bn , increasing to A$1.5bn at spot,
« Pre-tax IRR of 23% , increasing to 29% at spot,
« EBITDA margin (avg) of 44% , increasing to 49% at spot,
« Pre-Production capital costs of A$820M (incl contingency),
« Stage 1 payback of 2.7 years , reducing to 2.4 years at spot,
« Diversified revenue stream: Pd: 51%, Ni: 22%, Cu: 17% , Pt-Au-Co byproducts: 10%.
« ~50% of Resource remains unmined (~7.9Moz 3E, 450kt Ni, 250kt Cu, 46kt Co contained) below the modelled pit shell, providing exceptional upside potential to the above metrics.
« Gonneville set to become Australia’s first primary Platinum Group Metal (PGM) mine and 2[nd] largest nickel mine, producing an average of:
« 220kozpa of 3E precious metals (palladium, platinum and gold), plus;
« 7ktpa of nickel , 8ktpa of copper and 0.7ktpa of cobalt .
« Globally competitive 2[nd] quartile cost profile starting with US$50/oz 3E AISC net of byproduct credits in the first 3 years, trending to US$370/oz 3E avg over the 1.2x strip ratio open-pit phase:
« Lowest cost PGM mine in the western world and lowest cost of any undeveloped project.
« Exceptional margins expected to underpin significant, low-cost debt funding .

1 Base case byproduct price assumptions of Pt: US$1,300/oz, Au: US$2,900/oz, Co: US$39,000/t and AUD/USD: 0.65. 2 Spot prices Pd: US$1,500/oz, Pt: US$1,660/oz, Au: US$4,250/oz, Ni: US$14,900/t, Cu: US$12,050/t, Co: US$49,500/t, Cu conc TCRCs US$-40/t, US-4c/lb, Ni conc Ni payability 76%, sourced COMEX, LME, S&P Global 5 Dec 2025.

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Level 3, 46 Colin Street, West Perth WA 6005, Australia PO Box 428, West Perth WA 6872

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« Significant upside and leverage to higher commodity prices, accelerated expansions, optimisations, improved offtake terms and potential life extensions (beyond the scope of the PFS):
« Significant leverage to the palladium price, with a ~A$250M increase in pre-tax NPV8 and ~A$640M increase in cumulative pre-tax free cashflow, per US$100/oz increase in the longterm palladium price ( base case in bold ):

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« Stage 2 expansion from 5Mtpa to 14Mtpa process throughput expected to be funded out of post-financed cashflows – potential for acceleration or upsizing of the expansion according to macro-economic conditions during Stage 1 operations.
« Long term life extension potential through larger open-pit and/or large-scale underground operation from 2054+.
« High-grade mineralisation has been intersected ~900m beyond the limit of the Resource, which highlights the exceptional exploration upside of the Project.
« A saleable iron concentrate (predominantly magnetite) has been produced in testwork, providing a potential additional revenue stream in Stage 2 from 2034+.
« Potential for alternative downstream nickel processing customers from the pCAM industry which could provide improved offtake terms relative to existing nickel smelters.
« Maiden Ore Reserve[3] defined of 260Mt @ 0.86g/t 3E, 0.16% Ni, 0.098% Cu, 0.017% Co, containing 7.1Moz 3E, 400kt Ni, 250kt Cu, 43kt Co, which is limited by infill drilling.
« Fundable, executable and scalable two-stage development plan on 100% Chalice-owned farmland, proximal to major infrastructure and a residential workforce, with ‘ Major’ and ‘Strategic’ Project Status from WA/Commonwealth Governments:
« Project significantly de-risked with an investment of ~$240M by Chalice to date , including >1,200 resource definition drill holes, 33 dedicated metallurgical drill holes and over 1,400 metallurgical tests, 58 geotech-logged drill holes, extensive engineering trade-off studies, preliminary marketing discussions with smelters, active stakeholder engagement since the

3 Ore Reserves are reported at reserve prices of Pd: US$1,050/oz, Pt: US$1,000/oz, Au: US$2,200/oz, Ni: US$16,500/t, Cu: US$9,000/t, Co: US$30,000/t, AUD/USD: 0.65. Refer to JORC Table 1 for full details.

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discovery in 2020, acquisition of ~26km[2] of freehold title, significant environmental offset preparation and progression of environmental approvals in parallel to studies.

« Project materially improved and simplified over the last two years, with reduced upfront costs and risks, enhanced margins and financial metrics.
« Conventional upstream processing technologies utilised (Cu-Ni sequential flotation and precious metal leaching) to produce saleable smelter grade copper and nickel concentrates and precious metal doré.
« Gonneville will become a strategic, large-scale, western critical minerals mine :
« The first ever primary PGM mine in Australia , which helps diversify global supply (~93% of current PGM supply from Russia, South Africa and Zimbabwe).
« The critical minerals to be produced at Gonneville are essential to the auto sector (electric, hybrid and internal combustion engine vehicles), the defence sector (high performance materials and electronics), data centres (semiconductors and electrical components), as well as many rapidly growing decarbonisation applications .
« Project is strongly aligned to western government policy directives and helps address the critical minerals dominance of China, Russia and South Africa.
« Chalice is active in the local community and has completed three Local Voices surveys which indicate strong community support for the future mine and the economic development:
« Primarily a residential workforce expected due to proximity to Perth metro area – highly attractive long-term employment and business opportunities for the region.
« 1,200 FTE construction jobs in Stage 1 (2028-2029) and Stage 2 (2033)
« 500 FTE operations jobs for 23+ years (2029+)
« >A$1.5bn in direct royalties and taxes to state and federal governments.
« Chalice has engaged, early, actively and transparently to build respectful and collaborative relationships with stakeholders.
« The delivery of the Biodiversity Strategy and goals have commenced through the establishment of on-site restoration projects, with significant investment already made into environmental offsets .
« FID targeted for H1 CY28 – Chalice is fully funded to this milestone with A$76 2F [4] million in cash and listed investments – PFS completion allows commencement of the Feasibility Study, next stage of regulatory approvals and offtake/financing discussions.

Cautionary statement : The production targets disclosed in the PFS are based predominately on Measured (1%) and Indicated (93%) Mineral Resources. A small proportion of Inferred Resources (6%) have also been included. There is a low level of geological confidence associated with Inferred Mineral Resources and there is no certainty that further exploration work will result in the determination of further Measured and/or Indicated Mineral Resources or that the production targets associated with the Inferred Mineral Resources will be realised.

Chalice Managing Director & CEO, Alex Dorsch, said: “ The Pre-Feasibility Study is a major milestone for Chalice, and highlights our plan to develop a major critical minerals mine which, due to its scale, longevity and location represents a compelling, unique and strategic opportunity in the sector.

“The Study demonstrates that Gonneville is viable and will generate solid returns, even at bottom of the cycle prices. The Project has industry leading cost competitiveness, while will also deliver substantial social and economic benefits for the region, WA and Australia over a multi-decade life.

4 Includes ~$11M of listed investments at 30 September 2025.

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“The Project development plan has been dramatically simplified, with a two-stage open-pit and conventional concentrator-leach process flowsheet, which with recent improvements and optimisations in the PFS have underpinned a low-risk and financially strong study.

“The large, diverse production profile of the mine, together with its scalability, points to further upside particularly in volatile commodity price environments driven by the extreme scarcity of PGMs and the lack of large-scale base metal development projects worldwide. A range of enhancements and upside opportunities will continue to be developed in the Feasibility Study.

“You only need to look at similar large scale, low-grade deposits like Boddington and Cadia (owned by Newmont) to see how operations of this type can evolve and become larger over time.

“ The metals to be produced are currently dominated by Russia, China, South Africa and Indonesia. Gonneville will help address this dominance and provide a secure and reliable supply chain from a new independent producer in the western world.

We look forward to further formal discussions with smelters around offtake of our nickel and copper concentrates. We have already built strong relationships with ideal smelter customers and have seen exceptional interest levels in offtake and associated financing.

“We have also progressed the PFS with the help of Mitsubishi Corporation, who have provided technical and marketing guidance which has been highly valuable. We would like to thank Mitsubishi for their support and input to date, and we look forward to working closely with them as we progress into offtake and financing discussions.“

“I would like to take this opportunity to thank the entire project team and our consultants who have crafted a simple staged development plan which now provides a clear pathway to advance the Project towards development.

Chalice Chief Operating Officer, Dan Brearley, said: “Gonneville is a world-class, generational scale and iconic WA project that we are absolutely committed to developing in a sustainable and responsible manner. We will develop the Project with industry-leading execution discipline, environmental, community and cultural heritage management practices, ensuring long term positive impacts for local communities and regional stakeholders, because it is the right thing to do.

“ We now will progress the Feasibility Study, the next stage of regulatory approvals and continue to de-risk the Project through engineering. It’s a very exciting time for our team and we have come a long way in just over 5 years since our discovery.

“We have a fantastic in-house project team assembled, with strong support from our various service providers. I would also like to take the opportunity to thank them for the great work to date. We have all the right elements to now build a large project on the outskirts of Perth and we are moving up a gear to deliver it on time and on budget.”

Investor Conference Call

Chalice Mining Limited (“Chalice” or “the Company”, ASX: CHN) will host a live investor teleconference and webcast today commencing at 8.00am AWST/11.00am AEDT, Monday 8 December 2025 .

https://loghic.eventsair.com/375514/718326/Site/Register

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Project overview

Chalice Mining Limited (“Chalice” or “the Company”, ASX: CHN) is pleased to report the PreFeasibility Study (“PFS” or “Study”) for the 100%-owned Gonneville Project (“Gonneville”, the “Project”).

The Gonneville Project is located on Chalice-owned farmland (the “Mine Development Area”), ~70km north-east of Perth in Western Australia (Figure 1).

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Figure 1. Gonneville Project location.

The greenfield Project was staked in early 2018 as part of Chalice’s global search for high-potential nickel sulphide exploration opportunities. Limited exploration work had been conducted in the area prior to Chalice’s staking in 2018, owing to the lack of outcropping geology and perception of low prospectivity.

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A shallow, tier-1 scale polymetallic Resource was discovered by Chalice’s geologists in early 2020. The palladium-nickel-copper dominated Resource is one of the largest of its type in the western world and is one of the few amenable to open-pit mining.

The Resource hosts a rare mix of critical and strategic minerals, such as palladium, platinum, nickel, copper and cobalt, which are vital inputs into the auto sector (electric, hybrid and internal combustion engine vehicles), the defence sector (high performance materials and electronics), data centres (semiconductors and electrical components), as well as many rapidly growing decarbonisation applications (Table 1).

Table 1. Gonneville Mineral Resource Estimate (Resource) 23 April 2024

Classification Mass	Grade	Contained Metal
Mt	3E (g/t) Ni (%) Cu (%) Co (%)	3E (Moz) Ni (kt) Cu (kt) Co (kt)
Measured 2.9	1.20 0.21 0.17 0.018	0.12 6.1 4.8 0.52
Indicated 400	0.79 0.15 0.087 0.015	10 610 370 65
Inferred 250	0.80 0.15 0.076 0.014	6.4 370 200 37
Total 660	0.79 0.15 0.083 0.015	17 960 540 96

Resources reported above a pit constrained cut-off of A$25/t NSR and underground MSO cut-off of A$110/t NSR (refer to ASX Announcement 23 April 2024 for details of cut-off approach and assumptions). Note some numerical differences may occur due to rounding to 2 significant figures. 3E = Pd+Pt+Au at an approximate ratio of 4.5:1:0.15. The Resource underpinning the production targets in the Study has been prepared by a Competent Person and reported in accordance with the requirements of the JORC Code (2012).

The PFS describes a two-stage, open-pit critical minerals mine and process plant development which is predicted to become a large-scale producer of palladium, nickel and copper (co-products) over a modelled open-pit life of 23yrs, with valuable byproducts from cobalt, platinum and gold. The PFS has robust financial metrics which underpin the development of the Project at conservative longterm price assumptions.

The Study is based on the updated open-pit portion of the Gonneville Resource only and does not include an assessment of future underground mining nor extensions to mineralised zones beyond the Resource which have already been defined through step-out drilling.

The PFS mining inventory and mine life is limited by conservative mine design parameters rather than being constrained by Resources/drilling. The modelled open-pit exploits only ~50% of the current Resource, which remains open down-dip and to the north. High-grade mineralisation has been proven to extend ~900m beyond the limit of the Resource, which highlights the exceptional life extension upside of the Project.

The PFS development plan is materially different to previous project studies, with a two-stage development, a simplified flowsheet and design/optimisations based on a conservative, bottom of the cycle commodity price environment.

The PFS outlines a maiden Ore Reserve for the Project, that is limited to the open pit, Measured and Indicated portion of the Resource which has demonstrated economic viability (Table 2). Further conversion of Resources to Reserves is possible with infill drilling during operations.

Table 2. Gonneville Ore Reserve Estimate (Reserve).

Classification Mass	Grade	Contained Metal
Mt	3E (g/t) Ni (%) Cu (%) Co (%)	3E (Moz) Ni (kt) Cu (kt) Co (kt)
Proved 2.5	1.40 0.22 0.18 0.018	0.11 5.4 4.4 0.45
Probable 260	0.85 0.16 0.098 0.017	7.1 400 250 43
Total 260	0.86 0.16 0.098 0.017	7.1 400 260 43

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Ore Reserves are reported at reserve prices of Pd: US$1,050/oz, Pt: US$1,000/oz, Au: US$2,200/oz, Ni: US$16,500/t, Cu: US$9,000/t, Co: US$30,000/t, AUD/USD: 0.65. Refer to JORC Tables for full details. Note some numerical differences may occur due to rounding to 2 significant figures. The Reserve has been prepared by a Competent Person and reported in accordance with the requirements of the JORC Code (2012).

The Project is favourably located, with access to established road, rail, port and high-voltage power infrastructure nearby, plus access to a significant ‘residential’ mining workforce in the Perth surrounds.

In 2024, the Western Australian and Commonwealth Governments awarded ‘Strategic Project’ and ‘Major Project’ status to the Project, recognising its scale and strategic importance to the development of Australia’s critical minerals industry.

The Gonneville Project is expected to directly create around 1,200 jobs during peak construction and around 500 jobs per year in operation. These jobs will be particularly attractive given their proximity to Perth and the lifestyle values of the surrounding region.

Chalice recognises the need to develop the Gonneville Project sustainably, with a commitment to responsible environmental, social and cultural heritage management, and contribution to local economic development. Chalice is committed to rigorous standards and governing frameworks to ensure responsible environmental practices are followed in all our activities.

Commencing in 2020, Chalice progressively invested ~$50M to acquire a ~26km[2] package of freehold land, which covers the proposed mine development area. These acquisitions significantly de-risked the Project by providing certainty on tenure and provide a buffer to the limited neighbouring and biodiversity offset land properties.

Recognising the sensitivities of the area, Chalice has deliberately constrained the Project to Chaliceowned farmland. This land is already approximately 56% cleared from previous agriculture use. Developing a mine will have no material environmental impacts on neighbouring conservation areas.

Given the strategic and economic attractiveness of the Project, Chalice has formed the view that there is a reasonable basis to believe that requisite future funding for development of the Project will be available when required via a combination of both debt and equity. Informal discussions have commenced with potential financiers, indicating strong interest in the Project.

The Project has been significantly de-risked, with an investment of ~$240M by Chalice since the discovery in 2020. The Company is continuing to progress regulatory approvals, remaining studies, offtake and financing of the project ahead of a targeted Final Investment Decision (FID) in H1 CY28.

Key study outcomes and metrics

Production target

The PFS has outlined an outstanding opportunity to create a new long-life, low-cost, critical minerals mine in Western Australia, with significant upside and large-scale production profile. Gonneville is set to become the only primary PGM mine and the second largest nickel mine in Australia:

« Stage 1 (Years 1 to 4) : ~151koz 3E, 3.2kt Ni, 5.2kt Cu, 0.3kt Co per annum
« Stage 2 (Years 5 to 23) : ~238koz 3E, 7.7kt Ni, 8.7kt Cu, 0.7kt Co per annum

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Figure 2. Gonneville 3E precious metal production profile (koz, recovered).

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Figure 3. Gonneville base metal production profile (kt, recovered).

Financial return metrics

The PFS highlights the initial 23-year, two-stage open-pit phase has robust financial metrics using longterm, real base case commodity price assumptions of Pd: US$1,300/oz, Ni: US$18,750/t, Cu: US$10,500/t, Pt: US$1,300/oz, Au: US$2,900/oz, Co: US$39,000/t, approximating the ~95th percentile of industry cost curves (Table 3).

Project level financial metrics are presented at the base case prices as well as approximate spot prices and concentrate offtake terms as of 5 December 2025. All figures are in real terms (2025 AUD) and are unleveraged.

Table 3. Gonneville Project Pre-Feasibility Study key financial metrics (open-pit phase only).

Key metric Unit	Base case5	Spot case6
Modelled open-pit life Years 23 >23
Cumulative gross revenue A$bn 16.7 18.1
Cumulative EBITDA A$bn 6.9 8.5
EBITDA margin % 44 49
Cumulative free cashflow (pre-tax) A$bn 4.7 6.2
Cumulative free cashflow (post-tax) A$bn 3.6 4.7
Annual operating cashflow (pre-tax) A$Mpa 280 340
Annual operating cashflow (post-tax) A$Mpa 230 270

5 Wood Mackenzie 2025 nickel and copper cost curves sourced 31 Oct 2025, 95th percentile of palladium cost curve is Sibanye Stillwater US PGM Operations (2025 AISC guidance US$1,320/oz 2E incl S45X credit) sourced 7 Nov 2025).

6 Spot prices Pd: US$1,500/oz, Pt: US$1,660/oz, Au: US$4,250/oz, Ni: US$14,900/t, Cu: US$12,050/t, Co: US$49,500/t, Cu conc TCRCs US$-40/t, US-4c/lb, Ni conc Ni payability 76%, sourced COMEX, LME, S&P Global 5 Dec 2025.

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Key metric Unit	Base case5	Spot case6
NPV8%(pre-tax) A$bn 1.4 2.0
NPV8%(post-tax) A$bn 1.0 1.5
IRR (pre-tax) % 23 29
IRR (post-tax) % 21 26
NPC8%development CapEx A$bn 1.3
Stage 1 payback (from 1stproduction) Years 2.7 2.4
Stage 2 payback (from Yr5) Years 2.5 2.0
All-in Sustaining Costs (AISC)7 US$/oz 3E 370 390

Note: values are rounded to 2 significant figures. EBITDA margin calculated as portion of Net Smelter Return. NPC development CapEx is the net present cost of both stages of development capital, discounted to FID.

If the base case or higher prices are sustained over the longer term, the mine life is expected to well exceed the PFS modelled open-pit phase of 23 years, as the mining inventory is constrained to conservative mine design prices rather than the Resource (only ~50% of the Resource exploited by the PFS open-pit phase). Given this, there is considerable upside to the PFS metrics through expansions and/or life extensions.

The maximum negative free cashflow during the Stage 1 development is ~A$820M, including contingency. The Project is expected to generate pre-tax cashflows of A$300Mpa in the first 3 years, A$310Mpa for the next 10 years and A$240Mpa in years 13-23, at base case prices (Figure 4).

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Figure 4. Cashflow profile over modelled open-pit phase (pre-tax, real).

The two-stage development plan reduces overall execution risk and allows for the efficient deployment of capital. Importantly, given the scale and nature of the Gonneville Resource, the ability to expand the scale of the operation and/or drop the cut-off grade in future years is retained, providing exceptional optionality and leverage to higher long term metal prices.

7 AISC per produced 3E ounce (Pd+Pt+Au), net of byproduct credits after payabilities from Ni, Cu, Co. AISC calculation aligned to the SFA Oxford methodology, which excludes royalties, to compare with PGM industry peers.

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Detailed metrics by stage

Detailed financial, cost and physical metrics are presented for Stage 1 (the first 4 years), as well as the entire 23 year modelled open-pit phase of the Project, at base case prices (Table 4). Stage 1 has considerably higher feed grade than future years and hence considerably higher margins. All figures are in real terms (2025 AUD) and are unleveraged.

Table 4. Gonneville Project Pre-Feasibility Study detailed metrics by stage (base case prices).

Metric	Unit	Stage 1 (Years 1 to 4)	Modelled open-pit life (Years 1 to 23)
Financial
Gross Revenue (avg)	A$Mpa	450	730
Net Smelter Return per tonne processed	A$/t	90	57
EBITDA (avg)	A$Mpa	260	300
EBITDA margin (avg)	%	61	44
Annual operating cashflow (pre-tax)	A$Mpa	260	280
Annual operating cashflow (post-tax)	A$Mpa	250	230
Capital Costs
Pre-Prod development CapEx (incl. contingency)	A$M	820
Stage 2 expansion CapEx (incl. contingency)	A$M		840
Sustaining CapEx	A$Mpa	7	26
Operating Costs (avg)
Mine site cash costs per tonne processed	A$/t	35	32
Mine site cash costs per 3E ounce produced	US$/oz 3E	720	1,130
+ Transport & Selling costs	US$/oz 3E	57	53
- By-product credits (Ni, Cu, Co, Fe)	US$/oz 3E	700	890
= Total cash costs per 3E ounce	US$/oz 3E	75	290
+ Sustaining costs	US$/oz 3E	32	76
= All-in Sustaining Costs (AISC) per 3E ounce	US$/oz 3E	110	370
PGM Industry Cost Curve Position	quartile	1st	2nd
(net of by-products)
Mining Physicals
Total ore mined (excl pre-prod mining)	Mt	27	280
Total waste mined (excl pre-prod mining)	Mt	31	330
Total material movement incl. reclaim (avg)	Mtpa	14	36
Strip ratio (avg)	x	1.1	1.2
Processing Physicals
Total mass processed	Mt	19	280
« Measured	%	9	0
« Indicated	%	91	94
« Inferred	%	0	6
3E (Pd+Pt+Au) grade (avg)	g/t	1.44	0.85
Nickel grade (avg)	%	0.15	0.15

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Metric	Unit	Stage 1 (Years 1 to 4)	Modelled open-pit life (Years 1 to 23)
Copper grade (avg)	%	0.14	0.092
Cobalt grade (avg)	%	0.014	0.015
Oxide processing throughput	Mtpa	1	1→0
Oxide modelled life	Years		9
Sulphide processing throughput	Mtpa	4	4→12→14
Sulphide modelled life	Years		23
Produced 3E (Pd+Pt+Au)	koz	600	5,100
Produced nickel	kt	13	160
Produced copper	kt	21	186
Produced cobalt	kt	1.1	15
Pd recovery (avg)	%	71	74
Pt recovery (avg)	%	42	31
Au recovery (avg)	%	88	83
Ni recovery (avg)	%	44	38
Cu recovery (avg)	%	77	72
Co recovery (avg)	%	42	37

Note: values are rounded to 2 significant figures. Gross Revenue is net of payabilities (as invoiced by offtakers)

Cost profile

All-in Sustaining Costs (AISC) are calculated per total 3E (Pd+Pt+Au) precious metal ounce, which is consistent with the PGM Industry approach, given the Project is primarily driven by precious metals revenues (~58%) at base case prices.

The AISC is intended to highlight the costs and margins of the operation per produced 3E ounce. AISC is calculated as:

𝐴𝐼𝑆𝐶 (𝑈𝑆$/𝑜𝑧 3𝐸) =[𝑂𝑝𝐸𝑥+ 𝑠𝑢𝑠𝑡 𝐶𝑎𝑝𝐸𝑥−(𝑁𝑖+ 𝐶𝑢+ 𝐶𝑜 𝑟𝑒𝑣𝑒𝑛𝑢𝑒𝑠 𝑎𝑓𝑡𝑒𝑟 𝑝𝑎𝑦𝑎𝑏𝑖𝑙𝑖𝑡𝑖𝑒𝑠)] 𝑃𝑑+ 𝑃𝑡+ 𝐴𝑢 𝑝𝑟𝑜𝑑𝑢𝑐𝑒𝑑

The annualised AISC for Gonneville is very low during the initial years of production (~US$30/oz 3E in first 3 years), due to the shallow nature of the Resource and high grades near surface. The AISC per 3E ounce is improved by the strong byproduct revenue generated from the production of nickel, copper and cobalt (~42% of revenues).

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Figure 5. Gross revenue split by commodity and product (after payabilities), avg.

The AISC progressively trends up to US$370/oz 3E over the modelled open-pit life, primarily due to lower overall feed grades over time and higher mining costs as the open-pit gets deeper ( Figure 6).

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Figure 6. Gonneville AISC cost vs Gonneville 3E basket price 6F [8] over modelled open-pit phase.

In all years, the AISC is significantly below the base case long-term basket price (~US$1,354oz 3E), and well below the 70[th] percentile of the PGM industry cost curve (~US$1,180/oz 4E in 2024), highlighting the profitability of the operation through the commodity price cycle, its diversified revenue stream and its global competitiveness.

The low costs, significant margins and long-life of the Project also support the possibility of servicing significant long-term debt. Capital intensity assessment / benchmarking has not been performed, primarily because there have been very limited PGM development projects executed recently.

Industry competitiveness

8 Gonneville 3E basket price the weighted average Pd, Pt, Au price after payabilities.

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Byproduct credits from nickel, copper, chrome, cobalt, iridium, ruthenium and other minor metals are offset against costs.

It is noted that Russian and South African mines are responsible for >85% of 4E production (based on 2024 production). These countries have significant political, financial and operational challenges and the potential for supply disruptions from these countries is considered significant.

Gonneville is modelled to be 2[nd] quartile on the current 4E industry cost curve, and the lowest cost producer of PGMs in the western world, based on 2024 industry total cash and sustaining costs net of byproduct credits (Figure 7).

Norilsk Nickel (Russia) occupies the entirety of the first quartile and has negative cash costs due to their high level of Ni-Cu-Co by-product credits. Most South African PGM mines have very limited base metal by-product credits and typically involve very deep, narrow, non-mechanised underground mining with relatively high operating costs and significant development/sustaining costs.

Gonneville’s attractive position on the cost curve highlights a robust and competitive asset that is modelled to be highly profitable through the commodity cycle. The next best peer in the industry has AISC of ~US$721/oz (Impala Canada Lac Des Iles operation in 20248F[9] ), over double the predicted Gonneville AISC and has since announced closure plans in mid 2027 (costs artificially low at the end of the mine plan).

9 SFA Oxford 2024 actual, sourced on 4 June 2025

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Figure 7. 2024 PGM industry all-in sustaining cost curve (net of byproduct credits) and Gonneville positioning 7F [10] .

10 Source: 2024 SFA (Oxford) Ltd actual collated costs and revenues used for 4E cost curve data in June 2025. The Gonneville AISC assumes average by-product prices of: Copper US$10,500/t, Nickel US$18,750/t, Co US$39,000/t. AISC calculation aligned to the SFA Oxford methodology which excludes royalties.

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Sensitivity analysis

Sensitivity analysis has been performed as part of the PFS, assessing the robustness of the initial 23year, two stage open-pit phase financial metrics to a range of long-term metals prices, exchange rates, operating costs and capital costs as per industry standard practice (Figure 8).

All sensitivity analysis performed is within the financial model, which ignores the inherent ability to adapt to changing macro-economic conditions in real-time during operations of a large-scale, long life bulk open-pit, through:

« Adapting the mine design / mine plan due to changes in economic cut-off (increasing or decreasing the feed grade to plant and/or overall mine inventory),
« Adapting the process plant to chase higher recoveries through higher reagent use and higher operating costs,
« Applying hedging strategies,
« Increasing plant throughput capacity or performing retrofit / adaptations to the process plant configuration.

Therefore, the sensitivity analysis is indicative and does not reflect the true financial implications of significant movement in underlying assumptions, which can only be gauged through detailed mine redesigns or plant re-optimisations.

For the purposes of the sensitivity analysis on foreign exchange rates below, it is assumed 50% of CapEx and 25% of OpEx are effectively incurred (but not necessarily denominated) in USD, with the balance incurred in AUD. Offtake terms and payabilities remain fixed in the sensitivity analysis and are not varied with movements in metals prices.

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Figure 8. Modelled open-pit phase pre-tax NPV8 sensitivity analysis.

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Palladium, nickel and copper are the major revenue contributors, with palladium having a poor correlation to both copper and nickel historically, which provides a degree of diversification of the revenue stream and robustness to fluctuations in prices.

Over the mid 2023 to mid 2025 period, the palladium price remained well below the marginal cost of production (~US$1,320/oz) 9F[11] and has only recently recovered to more sustainable levels.

On the flipside, history demonstrates that when palladium rises above ‘incentive price’ levels (prices which incentivise capital investment to generate new supply), it can remain elevated for an extended period, as there are very limited additional sources of supply. In this way, historically palladium has shown an extremely low level of supply elasticity and hence very cyclical price behaviour, which is expected to continue into the future.

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Figure 9. LBMA palladium price (blue), PFS mine design and PFS long term price, US$/oz.

Sensitivity of key financial metrics to fluctuations in long term palladium, nickel and copper pricing has been performed, which highlights the significant leverage to higher long terms prices and robustness of the metrics to levels below the marginal cost of supply (Table 5).

11 Sibanye Stillwater 2025 guidance for US PGM Operations (US$1,320/oz 2E AISC incl S45X credit), sourced 7 Nov 2025)

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Table 5. Pre-tax NPV and IRR sensitivity to long term Pd, Ni and Cu prices (real). Base case in bold.

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The analysis demonstrates the robustness of the financial metrics even at the ~70[th] percentile of palladium industry cost curve (the price level where only 70% of producing mines in the industry are profitable) and at the low-end range of long-term nickel and copper price forecasts from industry banks/brokers.

This implies that even in a scenario where there is a 30% drop in palladium demand from current levels, assuming no supply cost escalation above the rate of inflation, Gonneville is still a viable Project (>15% IRR) to finance and execute. Chalice considers this extreme scenario unlikely however, given:

« The robustness of palladium demand, particularly from internal combustion and hybrid vehicles and slowing adoption of battery electric vehicles, but also from growing applications in data centres (electronic components, semi-conductors (multilayer ceramic capacitors) and precious metal investment given its extreme scarcity;
« Increasing palladium loadings per vehicle over time as emissions standards become stricter, particularly in the developing world;
« The lack of substitutes, or at least readily available substitutes in palladium applications;
« Palladium demand is extremely inelastic – i.e. consumers are not sensitive to the price, in particular when considering the input cost of palladium in an average internal combustion or hybrid vehicle is currently US$100-200 per vehicle;
« A prolonged subdued price environment for the metal will incentivise new applications (e.g. replacing gold in electrical connector plating, hydrogen production and purification and new chemical / catalytic applications), thus increasing demand over time;
« The instability and challenging investment landscape of Russia, South Africa and Zimbabwe;
« The rapid rise in industry costs in South Africa (>10% p.a.) driven by ageing and deep underground mines, which puts upwards pressure on prices over the long term;
« Structural infrastructure issues, corruption, political instability and high levels of inflation in South Africa and Zimbabwe;

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« Lack of palladium deposits and economically viable development projects globally (supply is extremely inelastic); and,
« Lack of investment in recycled / secondary supply, particularly without a significant, sustained price incentive above current levels.
« Negligible investment was made into recycling or any form of new supply in the last period of sustained incentive prices in 2019-2023, which demonstrates the extreme inelasticity of supply.

Development plan overview

The development plan for the Project includes an open-pit mine, process plant and supporting infrastructure, constructed in two stages. Stage 1 is designed for the lowest initial capital cost, maximum rate of return and shortest capital payback period, while the Stage 2 expansion is designed for optimal strategic value, mine life and profitability through the price cycle :

« Stage 1 – 4 years of higher-grade and higher-margin open-pit mining, processing oxide at 1Mtpa and sulphide at 4Mtpa in parallel, through a conventional crush-grind-flotation-leach process plant.
« Stage 2 – from year 5 to year 23, a long-life, bulk open-pit mining phase, processing oxide at 1Mtpa and sulphide at 12Mtpa processing throughput rate. De-bottlenecking of the process plant is completed post oxide feed exhaustion in year 9 to allow for an ultimate 14Mtpa sulphide process throughput rate.

The staged development approach de-risks the project with efficient deployment of capital and ability to adapt future stages to learnings and macro-economic conditions. Timing of the Stage 2 expansion is selected to ensure capital payback of Stage 1 and sufficient de-risking of the process flowsheet, however this could be accelerated if macro-economic conditions incentivise. Regulatory approval applications will include both Stage 1 and Stage 2, with any further expansions or line extensions needing future amendments.

The Stage 1 process throughput of 5Mtpa combined oxide and sulphide feed was selected as the optimal case for the higher-grade starter pit, which balanced sufficient return on fixed capital, shortest payback period, within funding constraints and a commensurate manageable risk profile for implementation by Chalice.

Ultimate processing capacity of 14Mtpa of sulphide feed was selected based on long term macroeconomic assumptions, mining inventory, equipment sizing, process water and site footprint characteristics, to deliver optimal strategic value of the project over the longer term within credible financing constraints. It is expected that significant debt funding would be available to fund both Stage 1 and Stage 2 capital costs.

The timing and sizing of the Stage 2 expansion is flexible and provides optionality, with the investment decision for this expansion expected to be made separately to Stage 1 FID, in ~2031-2033. Macroeconomic conditions may incentivise an earlier (or later) expansion, which would be possible within the planned regulatory approvals process.

If macro-economic conditions did not incentivise the expansion, a similar mine plan would essentially be followed but over a longer modelled life (~55 years as opposed to 23 years). Given the PFS financial outcomes however, Chalice considers both Stage 1 and Stage 2 to be incentivised at macro conditions well below the base case assumptions.

The Study is based on the open-pit portion of the Gonneville Resource only and does not include a likely transition to large-scale underground mining of the existing Resource in future, nor does it consider likely extensions to the Resource which have already been defined through step-out drilling.

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The PFS development plan is materially different to previous project studies, with a two-stage development, a simplified flowsheet and design/optimisations based on conservative, bottom of the cycle commodity price environment.

Mining

The Gonneville Resource starts at surface and hence conventional open-pit truck-excavator mining methods are selected for operations. Conventional grade-control, drill-and-blast and load-and-haul techniques are assumed, along with standard mining support fleet, all operated by a mining contractor.

The final pit dimensions are 1.7km (strike) x 1.0km (width) x 0.45km (depth). The pit shells are artificially constrained in the North to Chalice-owned farmland, inclusive of a buffer.

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Figure 10. 3D view (looking ENE) of the fresh sulphide ore blocks by NSR within the modelled pit.

The mine plan assumes a total material moved (TMM) rate of 14Mtpa for Stage 1, increasing to a maximum TMM rate of 38Mtpa in Stage 2. The mine plan assumes a level of stockpiling and rehandling to optimise grade to the process plant. Low-grade stockpiles and mine waste will be stored proximal to the open-pit, with any mineralised waste encapsulated progressively over time. The mine plan has a very low strip ratio (waste:ore) of 1.2 over the modelled life.

Importantly, the mine plan and cut-off grade may be adapted over time according to prevailing macro-economic conditions, which are highly cyclical. This cyclicality and optionality is not considered in the PFS, which assumes flat long-term prices in real terms. There is considerable value however inherent in this operational flexibility to adapt to price cycles.

Process plant

The process plant will produce three saleable products, including copper-palladium-platinum-gold and nickel-cobalt-palladium-platinum smelter concentrates and palladium-platinum-gold doré, utilising industry standard processing techniques (Figure 11).

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Figure 11. Gonneville Project Process Flowsheet (simplified).

The superficial free dig oxide Resource is processed using a conventional sizing, scrubbing and grinding circuit, followed by blending with the sulphide feed into a precious metal resin-in-leach adsorption process to produce a Pd-Pt-Au doré.

The fresh rock sulphide Resource is processed using a conventional crushing and grinding circuit utilising a SAG-ball-IsaMill[TM] configuration, a sequential sulphide flotation concentrator to produce two concentrates: a Cu-Pd-Pt-Au concentrate and a Ni-Co-Pd-Pt concentrate. The remaining sulphide feed is then blended with the oxide feed into the leach process to produce a Pd-Pt-Au doré.

Flowsheet and plant parameters are based on over three years of metallurgical testwork and flowsheet development, with a >$15 million investment by Chalice to date. This work included >1,000 flotation tests, >400 leach tests and full mass balances on seven metallurgical composites, derived from 33 dedicated metallurgical drill holes. As such, process plant performance has been materially de-risked.

Product marketing and offtake

The products are considered industry standard and commercially attractive to a broad range of potential customers. The products are expected to be marketed and sold as follows:

« The ~20% Cu, 45-60g/t 3E (Pd+Pt+Au) concentrate is expected to be sold directly to copper smelters in Asia and/or Europe, where offtake terms are expected to be highly favourable based on indicative terms received to date. The copper concentrate is expected to have negligible deleterious elements.
« The ~8% Ni, 0.8% Co, 18-20g/t 3E concentrate is expected to be sold directly to nickel smelters or pre-cursor Cathode Active Material (“pCAM”) refineries in Asia, Europe or North America, where offtake terms are expected to be favourable based on indicative terms received to date. The nickel concentrate is expected to have negligible deleterious elements, with a minor penalty for MgO in lower grade in the later years of the mine plan.
« The Pd-Pt-Au doré is expected to be sold directly to a precious metal refinery, where a nominal refining charge will be payable.

It is assumed that the payable metals in the offtake products will be nickel, copper, cobalt, palladium, platinum and gold, however, the concentrates do contain iron, rhodium, iridium, silver and other minor critical minerals, and the recovery and potential payability of these metals continues to be further investigated.

Supporting infrastructure and workforce

In addition to mining and processing facilities, waste storage, offices, temporary accommodation for construction workforce, roads/parking, stores and maintenance facilities will be built on site (Figure 12).

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Figure 12. Gonneville MDA preliminary site layout.

The Tailings Storage Facility (TSF) will be constructed in stages, as a downstream, high density polyethylene (HDPE) lined, valley-fill method. The TSF will have sufficient storage for the entire openpit modelled life, with further expansions to be subject to new regulatory approvals.

Mining and processing facilities on site will be supported by new power and water infrastructure, including a solar-battery-diesel hybrid power facility. The mine site will be connected to the South West Interconnected System (SWIS) electricity network to source power, via a new ~27km monopole dual-circuit high voltage transmission line from Muchea. A Connection Agreement is in place with Western Power to progress scoping of this infrastructure.

Process water is to be supplied via a new ~63km pipeline to the Alkimos Wastewater Treatment Plant. A Letter of Intent (LOI) has been executed with Water Corporation in relation to the offtake of treated wastewater, which is currently being discharged into the ocean. The forecast volume of water supply available at Alkimos provides sufficient volume for the modelled open-pit life of the Project and is expected to increase over time with the expected expansion of the Perth metropolitan area.

Two potential water and power infrastructure corridors have been scoped with flora and fauna surveys ongoing, and heritage surveys planned in CY26. Government Trading Entities Western Power and Water Corporation continue to be engaged on cost and execution schedule for this infrastructure. Chalice continues to engage with the WA and Commonwealth Governments around potential common user infrastructure options and funding support.

Bulk copper and nickel concentrates are assumed to be trucked and exported via the Port of Bunbury in Stage 1. In Stage 2, concentrates are assumed to be trucked and exported via the planned new Kwinana Bulk Terminal Port.

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The construction workforce is assumed to be largely residential (locally based, commuting to site daily) with consideration of some temporary accommodation on site, while the operations workforce is assumed to be residential.

Regulatory approvals

Gonneville has received formal recognition from both the State and Commonwealth Governments with ‘Strategic Project’ status from the WA State Government and ‘Major Project’ status from the Commonwealth Government of Australia, recognising the strategic significance of the project. This formal recognition provides a level of prioritisation and streamlining of regulatory assessment by governments. These designations do not bypass or weaken environmental approvals processes.

The Project will require approvals under the WA Environmental Protection Act 1986 , WA Mining Act 1978 , and the Commonwealth Environment Protection and Biodiversity Conservation Act 1999 .

Chalice owns ~26km[2] of freehold land over and surrounding the 22km[2] Mine Development Area (MDA). The MDA is subject to exploration tenure granted under the WA Mining Act 1978 , comprising Exploration Licences E70/5118, E70/5119 and E70/5353.

To progress the Project, it is intended that portions of this exploration tenure coinciding with the MDA will ultimately be converted to Mining Lease(s). The area applied for under a Mining Lease will encompass the Gonneville mine footprint and all associated mining and processing facilities.

The infrastructure corridors for power and water pipeline required to support the mine development will be progressed for approval via a Miscellaneous Licence(s).

Extensive work has been undertaken by Chalice to develop environmental baselines and define the programme of environmental surveys and studies required to support formal environmental assessment during the study phase of the Project.

Formal referral of the Project to State and Commonwealth Governments was submitted in March 2024 which commenced the regulatory environmental approvals processes. It is anticipated that the Environmental Review Documents (ERDs) will be submitted in H2 CY26. Indicative approval timelines, which govern the overall project development timeline, are estimates only and not all steps in the approvals process are subject to statutory timeframes and could vary to those anticipated.

Community

From the discovery and early development of the Project in 2020, Chalice has recognised that the local community are our most important stakeholders, and effective community engagement is critical to the success of the Project.

Chalice has proactive, regular and transparent engagement with key stakeholders across the local community, indigenous landowner groups and all levels of government. At all levels of the Company, Chalice seeks to engage with a genuine, responsive, and approachable style to establish the Company as a trusted organisation.

This engagement has been reflected in the results of three Local Voices surveys conducted since 2023, which have all indicated high levels of support for the Project.

Chalice has invested more than ~$11M into the local community via local spend and direct contractors. The mine is expected to generate 1,200 FTE construction jobs and 500 FTE operations jobs and contribute more than $1.5 billion in direct royalties and taxes to the state and federal governments. The workforce will be largely residential, based in and around the surrounding areas of Perth, making it a highly sought after location for workers in the mining industry.

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The Project is located in the Whadjuk Indigenous Land Use Agreement (ILUA) area, signed as part of the historic 2021 South West Native Title Settlement between the Noongar people and the Western Australian Government.

Engagement with Whadjuk has been ongoing since mid-2021 through cultural awareness and cultural safety training, including On-Country visits, involvement in environmental surveys as well as the completion of cultural heritage surveys covering all Chalice-owned farmland in and around the Project site. The completed cultural heritage survey report identifies no impediments to the advancement of the Project.

Chalice has received a positive reception from Whadjuk People and is committed to further developing this important relationship as the Project matures and delivers direct benefits to the group.

Environment

Chalice is committed to rigorous standards and governance frameworks to ensure responsible environmental practices are followed in all our activities.

Decisions on design, construction and operation are guided by the mitigation hierarchy to manage and minimise our impacts on the environment. This hierarchy of Avoid, Minimise, Rehabilitate and Offset, underpins all aspects of our environmental management and reflects our commitment to responsible mining practices.

The Tailings Storage Facility (TSF) and other waste management facilities and practices have been designed to comply with a zero-discharge site policy i.e. no pollutants discharged to the external environment. The TSF design complies with relevant WA state guidance and code of practices, national ANCOLD 10F[12] guidelines and the Global Industry Standard on Tailings Management (GISTM 2020).

We fundamentally believe that mining can be undertaken sustainably and responsibly, and that the Project can co-exist with conservation and community values. Further information regarding our approach to Sustainability can be found in the 2025 Sustainability Report, part of the Annual Report (refer to ASX Announcement on 26 September 2025).

Our commitment to biodiversity is highlighted in our Gonneville Biodiversity Strategy that spans the life of the mine and beyond, to achieve two key goals:

« To ensure science-based no net loss of species or habitat as a result of any mining operations.
« To strive towards a net positive legacy for significant species and our local community.

The Biodiversity Strategy and goals will be delivered through on-the-ground restoration projects that increase habitat availability and connect remnant areas of habitat on farmland and adjacent areas of the conservation estate that are currently fragmented.

Approximately 400ha of Chalice-owned land adjacent to the MDA (Figure 13) have been designated as Biodiversity Offset areas. A detailed implementation plan has been developed, and work has commenced with the establishment of a Pilot Restoration Area (Figure 14), and research partnerships focusing on Chuditch and Black Cockatoo species, the key threatened fauna species whose protection and management is critical to the success against the goal of ‘no net loss of species’.

The findings of these studies will inform both the appropriate restoration responses and control management actions for the restoration areas.

12 Australian National Committee on Large Dams

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Figure 13. Project Mine Development Area and planned offset areas.

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Figure 14. Pilot restoration area establishment.

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Project funding

Chalice has been engaging with a number of potential finance partners from both public and private markets during the PFS. Governments, both Australian and international, have indicated a strong appetite to support funding of critical minerals projects like Gonneville in western countries. Palladium, platinum, nickel and cobalt are all considered critical minerals by the Australian Government, whilst copper is considered a ‘strategic mineral’.

Funding is expected to be sourced from range of partners including:

« Western Australian State Government sponsored initiatives
« Australian Federal Government sponsored initiatives
« International government sponsored initiatives
« Offtake partners
« Specialist ‘green’ finance providers
« Commercial banks

A detailed funding plan will be developed during the Feasibility Study. Chalice has formed the view that there is a reasonable basis to believe that requisite future funding for development of the Project will be available when required. The grounds on which this reasonable basis is established includes:

« Australian and international governments, from Organisation for Economic Co-operation and Development (OECD) countries, have a strong appetite to support large scale critical minerals projects.
« Export Credit Agencies (ECAs) from major OECD nations, including Australia, Canada, Germany, Japan, and the United States, are showing increasing interest in financing projects such as Gonneville as part of broader national strategies to enhance supply chain resilience.
« ECAs can provide long-term, low-cost debt which bolsters project viability and profile, often catalysing equity participation from sovereign wealth funds, development finance institutions, and strategic investors seeking de-risked exposure to critical minerals.
« The signing of a non-binding Memorandum of Understanding (MoU) with Mitsubishi Corporation in 2024 highlights clear potential for strategic partnership in the development of the Project. This collaboration demonstrates growing global interest in securing reliable sources of critical minerals and reinforces Gonneville’s appeal to world-class industrial counterparties.
« Chalice has a current market capitalisation of approximately A$680 million (at 5 December 2025) and no debt.
« The Company has a strong track record of successfully raising equity funds in a prudent and disciplined manner when required to further the exploration and development of the Project.
« The Chalice Board and management team has experience in mine development, financing and operations in the resources industry.
« Chalice owns 100% of the Gonneville Project and there are no historical financing mechanisms (e.g. royalty, stream, etc) encumbering its development.
« The Project is located in Western Australia, which is considered one of the lowest-risk, most stable and attractive jurisdictions globally for mining.
« The Project has an initial modelled open-pit life of 23 years.
« Project economic viability has been established at bottom of the cycle commodity prices, with the PFS demonstrating a highly competitive 2[nd] quartile position on the PGM industry cost curve

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and an unleveraged payback period of under three years – there is considerable financial capacity to cover long term debt repayments.

Chalice has $76M in cash and listed investments (at end September 2025) and is fully funded through to targeted FID in H1 CY28. Project financing is expected to be secured following completion of the Feasibility Study, which is targeted for H1 CY27.

Development timeline and next steps

The PFS has demonstrated that the Project is technically and commercially viable, and hence Chalice is now progressing the development plan into a Feasibility Study (FS). The FS will involve optimising the design and undertaking detailed engineering to prepare the Project for a Final Investment Decision (FID) on Stage 1, targeted in H1 CY28.

The Company is targeting submission of the Environmental Review Documents (ERDs) in H2 CY26 using the PFS development plan as the basis for the submission. Importantly, the approval scope will consider the full scale and long-term impacts of the Project, so there is scope to adjust the staging of construction according to macro-economic conditions.

Offtake negotiations for copper and nickel concentrates will commence in CY26, with the aim of securing foundational customers for these products, whilst maintaining flexibility and optionality for as long as possible. Offtake discussions could potentially include linked project financing, as a favourable source of capital and mechanism for alignment with downstream partners.

An FID is expected to be made, subject to the finalisation of all key activities:

« Feasibility Study completed H1 CY27 « Offtake agreements executed H2 CY27 « Funding sourced H1 CY28
« Major environmental approvals H1 CY28

Following FID, a 1.5 to 2-year engineering and construction phase is expected, resulting in first production in early 2030 (Figure 15).

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Figure 15. Gonneville Project overall development schedule.

Further detailed information can be found in the following Executive Summary of the Pre-Feasibility Study Report.

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Authorised for release by the Board of Directors.

For further information, please visit www.chalicemining.com, or contact:

Corporate Enquiries

Ben Goldbloom GM Corporate Development Chalice Mining Limited +61 8 9322 3960 [email protected]

Media Enquiries

Follow our communications

Nicholas Read LinkedIn: chalice-mining Principal and Managing Director Twitter: @chalicemining Read Corporate Investor Relations +61 8 9388 1474 [email protected]

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Pre-Feasibility Study

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Gonneville Pd-Ni-Cu Project

1.		Project location and history ................................................................................................... 5
	1.1	Mine Development Area ....................................................................................................... 7
	1.2	Landform and topography ................................................................................................... 7
	1.3	Flora and fauna ...................................................................................................................... 7
	1.4	Surface water and ground water ......................................................................................... 8
	1.5	Climate ..................................................................................................................................... 8
2.		Development plan overview................................................................................................. 9
	2.1	Mining ..................................................................................................................................... 10
	2.2	Process plant ......................................................................................................................... 10
	2.3	Product marketing and offtake .......................................................................................... 11
	2.4	Supporting infrastructure and workforce ........................................................................... 11
3.		Geology and Mineral Resources ........................................................................................ 13
	3.1	Geology ................................................................................................................................. 13
	3.2	Mineralisation ........................................................................................................................ 13
	3.3	Mineral Resources ................................................................................................................. 14
4.		Geotechnical ........................................................................................................................ 19
	4.1	Data collection ..................................................................................................................... 19
	4.2	Rock Mass Strength ............................................................................................................... 19
	4.3	Output data .......................................................................................................................... 20
	4.4	Pit slope design ..................................................................................................................... 21
	4.5	Stability analysis ..................................................................................................................... 22
5.		Mining and Ore Reserves ..................................................................................................... 22
	5.1	Dilution analysis ..................................................................................................................... 23
	5.2	Open-pit mine design .......................................................................................................... 23
	5.3	Open-pit mine schedule ...................................................................................................... 28
	5.4	Ore Reserve ........................................................................................................................... 32
6.		Metallurgy and processing .................................................................................................. 32
	6.1	Metallurgical testwork .......................................................................................................... 33
	6.1.1	Feed samples .......................................................................................................................... 33
	6.1.2	Process water samples ........................................................................................................... 36
	6.2	Mineralogy ............................................................................................................................. 36
	6.2.1	Sulphide ................................................................................................................................... 36
	6.2.2	Oxide ........................................................................................................................................ 37
	6.3	Process and flowsheet design ............................................................................................. 37
	6.3.1	Trade-off studies ..................................................................................................................... 40
	6.3.2	Oxide comminution ............................................................................................................... 42
	6.3.3	Sulphide comminution ........................................................................................................... 43
	6.3.4	Sulphide flotation concentration ......................................................................................... 43
	6.3.5	Leach ....................................................................................................................................... 46

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6.4	Geo-met and recovery modelling ..................................................................................... 48
6.4.1	Oxide recovery ....................................................................................................................... 48
6.4.2	Transitional recovery .............................................................................................................. 49
6.4.3	Fresh sulphide recovery ......................................................................................................... 50
7.	Tailings and waste management ....................................................................................... 50
7.1	Tailings Storage Facility design ............................................................................................ 51
8.	Infrastructure and logistics ................................................................................................... 52
8.1	Process water supply ............................................................................................................ 53
8.2	Power supply ......................................................................................................................... 54
8.3	Logistics .................................................................................................................................. 54
8.4	Non-process infrastructure (NPI) ......................................................................................... 56
9.	Tenure, approvals and stakeholder engagement ........................................................... 56
9.1	Tenure ..................................................................................................................................... 56
9.2	Environmental approvals ..................................................................................................... 57
9.3	Native Title, heritage and traditional owner participation .............................................. 60
9.4	Stakeholder engagement ................................................................................................... 60
9.4.1	Community investment.......................................................................................................... 61
10.	Development timeline and implementation ..................................................................... 62
11.	Cost estimates ....................................................................................................................... 63
11.1	Development capital expenditure (CapEx) estimates .................................................... 63
11.2	Sustaining capital expenditure (CapEx) estimates ........................................................... 64
11.3	Operating expenditure (OpEx) estimates .......................................................................... 65
12.	Product marketing and offtake .......................................................................................... 65
12.1	Palladium market overview ................................................................................................. 66
12.2	Cu-PGM concentrate .......................................................................................................... 67
12.3	Ni-Co-PGM concentrate...................................................................................................... 68
12.4	PGM doré ............................................................................................................................... 69
13.	Financial analysis................................................................................................................... 69
13.1	Key assumptions .................................................................................................................... 69
13.2	Financial return metrics ........................................................................................................ 71
13.3	Detailed metrics by stage .................................................................................................... 73
13.4	Cost profile ............................................................................................................................. 74
13.5	Industry competitiveness ...................................................................................................... 76
13.6	Sensitivity analysis .................................................................................................................. 78
13.7	Project funding ...................................................................................................................... 81
14.	Upside opportunities and risks ............................................................................................. 82
14.1	Additional iron byproduct upside ....................................................................................... 82
14.1.1	Testwork ................................................................................................................................... 82
14.1.2	Logistics and marketing ......................................................................................................... 82
14.1.3	Economic analysis .................................................................................................................. 83
14.2	Resource/mining upside ...................................................................................................... 83

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14.3	Processing upside ................................................................................................................. 85
14.4	Commercial upside .............................................................................................................. 86
14.5	Project risks ............................................................................................................................. 86
15.	Study team ............................................................................................................................ 88
16.	Appendices ........................................................................................................................... 90
A-1	Metallurgical Recovery Technical Data ............................................................................ 90
A-1-1	Sulphide ore mass balances ................................................................................................. 90
A-2	Competent Person Statement ............................................................................................ 94
A-2-1	Mining and Reserves .............................................................................................................. 94
A-2-2	Metallurgy ................................................................................................................................ 94
A-2-3	Exploration Results .................................................................................................................. 94
A-2-4	Mineral Resources .................................................................................................................. 94
A-2-5	Resource estimation methodology ...................................................................................... 95
A-2-6	Forward Looking Statements ................................................................................................ 96
A-2-7	Non-IFRS Financial Measures ................................................................................................. 96
A-3	JORC Tables........................................................................................................................... 97
A-3-1	Section 1: Sampling Techniques and Data ......................................................................... 97
A-3-2	Section 2: Reporting of Exploration Results........................................................................ 102
A-3-3	Section 3: Estimation and Reporting of Mineral Resources ............................................. 104
A-3-4	Section 4: Estimation and Reporting of Ore Reserves ...................................................... 119

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1. Project location and history

The 100%-owned Gonneville Palladium-Nickel-Copper Project (“Project” or “Gonneville” is located on Chalice-owned farmland, ~70km north-east of Perth in Western Australia (Figure 1).

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Figure 1. Gonneville Project location map.

The greenfield Project was staked in early 2018 as part of Chalice’s global search for high-potential nickel sulphide exploration opportunities.

A shallow, tier-1 scale polymetallic critical minerals Resource was discovered by Chalice’s geologists in early 2020. The palladium-nickel-copper dominated Resource is one of the largest of its type in the western world and is one of the few amenable to open-pit mining.

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The Resource hosts a rare mix of critical and strategic minerals such as palladium, platinum, nickel, copper and cobalt. Palladium and platinum are some of the rarest elements found on Earth, with an average crustal abundance below that of gold and most other elements[1] .

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Figure 2. Abundance (atom fraction) of the chemical elements in Earth's upper continental crust as a function of atomic number.

The minerals to be produced at Gonneville are critical to the auto sector (electric, hybrid and internal combustion engine vehicles), the defence sector (high performance materials and electronics), data centres (semiconductors and electrical components), as well as many rapidly growing decarbonisation applications.

In addition, the Project is strongly aligned to western government policy directives in securing new sources of critical minerals supply and will help address the critical minerals dominance of China, Russia and South Africa.

The maiden Resource was defined in 2021 and has subsequently been updated three times, with increasing size and understanding of the geological and geo-metallurgical model.

Chalice’s study approach has been to define and de-risk a project which balances value, optionality and risk. Given the size and nature of the Resource, a staged development plan was viewed as the preferred utilisation and deployment of capital and the best balance of value and risk.

A Scoping Study was completed in 2023 which investigated several open-pit mine plans utilising a concentrator-hydromet-leach base-metal flowsheet. Subsequent development testwork in 2024 resulted in a vast improvement in flotation performance and the ability to produce saleable copper and nickel concentrates, which dramatically simplified and improved project viability.

A Pre-Feasibility Study (“PFS” or “Study”) commenced in 2023, to evaluate the two-stage development of a long-life critical minerals mine and associated process plant. The Study is based on the updated open-pit portion of the Gonneville Resource only and does not include extensions

1 "Abundance of Elements in the Earth's Crust and in the Sea", CRC Handbook of Chemistry and Physics, 97th edition (2016–2017)

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to mineralised zones which have already been defined through step-out drilling. Downstream processing options have not been included in the PFS but will be considered in the future as commodity markets evolve.

In 2024, Chalice referred the Project to the Western Australia (WA) and Commonwealth regulators for major environment approvals. In the same year, the WA and Commonwealth Governments awarded ‘Strategic Project’ and ‘Major Project’ status to the Project, recognising its scale and strategic importance to the development of Australia’s critical minerals industry.

Commencing in 2020, Chalice progressively invested ~$50 million to acquire a ~26km[2] package of freehold land which covers the Resource and surrounding areas. These acquisitions significantly derisked the Project by providing certainty on tenure and affording a buffer to the limited neighbouring properties and biodiversity offset land.

1.1 Mine Development Area

The Mine Development Area (MDA) is located in the Wheatbelt Region of WA, within the Shire of Toodyay - which has a population of approximately 4,600 (2021 census).

The Project is favourably located just over a 1-hour drive from Perth’s central business district (CBD), with access to established road, rail, port and high-voltage power infrastructure nearby, along with access to a significant and highly skilled ‘residential’ mining workforce in the Perth surrounds.

The mine is adjacent to the Julimar State Forest to the north, the Moondyne Nature Reserve to the south, and rural properties to the west and east. The dominant land use across these areas is agriculture (grazing and cropping) and rural lifestyle properties.

The supporting infrastructure corridors collectively intersect with three local government areas, the City of Wanneroo, the Shire of Chittering and the Shire of Toodyay.

Both the mine and associated infrastructure corridors lie within the Whadjuk People Indigenous Land Use Agreement Area (ILUA, WI2017/015).

1.2 Landform and topography

The MDA lies on the Darling Plateau, which comprises an undulating landform from 155m to 275m Above Mean Sea Level (AMSL) of lateritic regolith over Archean age predominantly granitic rocks.

The surficial geology of the area comprises laterite, with sand sheets overlying laterite and alluvium in headwaters of drainage areas, and granitic rocks (Age) exposed or at shallow depth in drainage areas on the south-west and east sides. The infrastructure corridor options run east to west and intersect both the Swan Coastal Plain and the Darling Plateau.

1.3 Flora and fauna

The Project is located within the Jarrah Forest bioregion, as described by the Interim Biogeographic Regionalisation for Australia (IBRA). The Jarrah Forest bioregion is classified into two subregions, Northern Jarrah Forest and Southern Jarrah Forest, with the MDA located within the Northern Jarrah Forrest subregion.

The Northern Jarrah Forest subregion is characterised by jarrah – marri forest on laterite gravels in the west, with bullich and blackbutt in the valleys, grading to wandoo – marri woodlands on clayey soils in the east, with powder bark on breakaways. There are also extensive, but localised, sand sheets with banksia low woodlands, and heath is found on granite rocks and as a common understory of forests and woodlands in the north and east.

The MDA is located across several largely cleared farming properties and covers approximately 2,240 hectares. These have been subject to several flora and vegetation surveys undertaken between 2020 and 2024, by Biologic Environmental Survey Pty Ltd (Biologic).

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A total of 16 vegetation types has been described and mapped across the MDA, representative of the following broad landforms; Hills, Deep Sands, Valleys, Drainage lines and Wetlands.

The land has been previously cleared for agriculture and, as such, the vegetation condition[2] varies. Most of the MDA is ‘Degraded’ or ‘Completely Degraded’ with signs of impacts associated with livestock, including trampling, grazing and clearing. The areas mapped as ‘Completely Degraded’ include farmlands and planted areas (~56%).

Biologic conducted Terrestrial Fauna surveys across the MDA between 2020 and 2024 to document vertebrate fauna values to inform the environmental impact assessment. Various survey methods including habitat assessment and mapping, cage traps, GPS tracking, camera traps, ultrasonic recording, targeted searches and opportunistic records were used to document fauna species within the area. Five conservation significant fauna species were recorded across the MDA; Chuditch, Carnaby’s black cockatoo, Forest red-tailed black cockatoo, Quenda and Western brush wallaby. A desktop assessment of the MDA, which included a 20km buffer found that these fauna species are typical of what is usually found in the Northern Jarrah Forest subregion.

1.4 Surface water and ground water

The MDA is located within the catchment of the Avon River, with the western portion drained by two un-named tributaries of the Brockman River and the eastern portion draining into the Julimar Brook. The Brockman River and Julimar Brook flow to the south through rural properties and discharge into the Avon River.

Surface water flow within the MDA is seasonal, with no natural perennial water bodies present. Farm dams comprise artificial perennial and semi-perennial water bodies within the area.

The MDA is located within the western part of the Yilgarn Craton, on the Darling Plateau. The Yilgarn Craton is largely composed of granitic rocks of the Pre-Cambrian age Western Shield. The main aquifer systems found in the Yilgarn Craton relevant to the MDA are Quaternary-age deposits along modern drainage lines, weathered bedrock and fractured bedrock. Fresh, unfractured bedrock at the MDA is expected to have low potential to act as an aquifer.

Aquifer recharge is via rainfall infiltration. Most of the rainfall is lost by evaporation or runoff and only a very minor portion of rainfall infiltrates through the soil and recharges the groundwater. The groundwater table of the entire aquifer follows the regional topographic gradient and tends to come closer to the surface in valleys.

Groundwater investigations to date have yielded limited supply potential.

1.5 Climate

The climate of the region is classified by cool wet winters, and warm, relatively dry summers. Average annual rainfall for the Northern Jarrah Forest subregion ranges from 1300 millimetre (mm) on the scarp to approximately 700mm in the east and north. The nearby weather stations likely to accurately document the long-term average temperature and rainfall for the MDA are the Bureau of Meteorology’s (BOM) Pearce Royal Australian Air Force (RAAF) and Lower Chittering weather stations (station numbers 9053 and 9009, respectively), located approximately 20km and 8km to the southwest, respectively.

The hottest month for Pearce RAAF is January (mean maximum temperature 33.6°C) while the coolest month is July (mean maximum temperature 17.9°C), where night-time temperatures regularly fall below 10°C (the month of August has the lowest mean minimum temperature of 8.3°C) (length of record from 1937 – 2024)[3] . The average annual rainfall for Lower Chittering is 810.5 mm, with

2 Vegetation condition scale adapted from Keighery, 1994. Technical Guidance – Flora and Vegetation Surveys for Environmental Impact Assessment WA: EPA 2016

3 Bureau of Meteorology 2024. Climate Data Online - https://www.bom.gov.au/climate/data/index.shtml

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average monthly rainfall peaking from late autumn to early spring (May to September). The highest average monthly rainfall occurs in July (158.8mm), with the lowest occurring in January (11.8mm).

2. Development plan overview

« Stage 1 – 4 years of higher-grade and higher-margin open-pit mining, processing oxide at 1Mtpa and sulphide at 4Mtpa in parallel, through a conventional crush-grind-flotation-leach process plant.
« Stage 2 – from year 5 to year 23, a long-life, bulk open-pit mining phase, processing oxide at 1Mtpa and sulphide at 12Mtpa processing throughput rate. De-bottlenecking of the process plant is completed post oxide feed exhaustion in year 9 to allow for an ultimate 14Mtpa sulphide process throughput rate.

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2.1 Mining

The final pit dimensions are 1.7km (strike) x 1.0km (width) x 0.45km (depth). The pit shells are artificially constrained in the North to Chalice-owned farmland, inclusive of a buffer.

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Figure 3. 3D view (looking ENE) of the fresh sulphide ore blocks by NSR within the modelled pit.

2.2 Process plant

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Figure 4. Gonneville Project Process Flowsheet (simplified).

2.3 Product marketing and offtake

The products are considered industry standard and commercially attractive to a broad range of potential customers. The products are expected to be marketed and sold as follows:

« The ~20% Cu, 45-60g/t 3E (Pd+Pt+Au) concentrate is expected to be sold directly to copper smelters in Asia and/or Europe, where offtake terms are expected to be highly favourable based on indicative terms received to date. The copper concentrate is expected to have negligible deleterious elements.
« The ~8% Ni, 0.8% Co, 18-20g/t 3E concentrate is expected to be sold directly to nickel smelters or pre-cursor Cathode Active Material (“pCAM”) refineries in Asia, Europe or North America, where offtake terms are expected to be favourable based on indicative terms received to date. The nickel concentrate is expected to have negligible deleterious elements, with a minor penalty for MgO in lower grade in the later years of the mine plan.
« The Pd-Pt-Au doré is expected to be sold directly to a precious metal refinery, where a nominal refining charge will be payable.

2.4 Supporting infrastructure and workforce

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Figure 5. Gonneville MDA preliminary site layout.

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3. Geology and Mineral Resources

3.1 Geology

The Gonneville Deposit is located within a ~1.9km x 0.9km x >0.8km section of the Julimar Complex, known as the Gonneville Intrusion - which has a north-north-east strike, maximum thickness of approximately 650m, and a 45° west-north-west dip. The Gonneville Intrusion is composed predominantly of serpentinised olivine peridotite/harzburgite (serpentine-magnetite-amphibolechromite) with lesser intervals of pyroxenite (amphibole-chlorite), gabbro and leucogabbro (clinozoisite-amphibole) divided into a series of eight litho-geochemical domains.

The weathering profile in the area extends to approximately 30-40m below surface. A welldeveloped laterite and saprolite profile is present which contains elevated palladium grades from near-surface to a depth of approximately 25m. There is a narrow transition zone between the oxide and sulphide zones, which is generally <15m thick.

The litho-geochemical domains broadly parallel the strike and dip of the Gonneville Intrusion and are interpreted to represent discrete magma influxes and associated fractionation units. The intrusion is crosscut by a later granite body, which broadly parallels the dip and strike orientation of the maficultramafic package.

Crosscutting the entire intrusive package is a series of sub-vertical, north-east to north-west striking, dolerite dykes. Both the granite body and dolerite dykes are un-mineralised. A package of metasedimentary rocks surrounds the Gonneville Intrusion. Although texturally the intrusive rock-types within the complex are moderately well preserved, permitting the use of igneous terminology, all rock units have been replaced by mineral assemblages characteristic of upper greenschist to lower amphibolite facies metamorphism.

The Gonneville Intrusion is bounded to the west (hanging wall) by felsic gneiss/metasediment and to the east (footwall) by a succession comprising metasediments (sulphidic pelite) and amphibolite of uncertain parentage.

3.2 Mineralisation

Primary Pd-Pt-Ni-Cu-Co-Au sulphide mineralisation occurs principally within the ultramafic domains of the Gonneville Intrusion and, to a lesser extent in gabbro subunits. Mineralisation is present as subparallel sulphide-rich zones (>20% sulphides), typically 5-40m wide, that occur within broader intervals (~100-150m wide) of weakly disseminated sulphides. The orientation of the higher-grade mineralised sulphide zones suggests an association with the litho-chronological domains within the intrusion (Figure 6).

There are four typical sulphide mineralisation types recognised at Gonneville:

« Massive sulphides: >75% (by volume) sulphide;
« Matrix sulphides: 40% to 75% sulphide; also referred to as net-textured, typically occurs as interconnected pyrrhotite-pentlandite-chalcopyrite mineralisation with silicate gangue;
« Stringer sulphides: 10% to 75% sulphide. Stringer sulphide mineralisation is typically observed around faults or lithological contacts; and
« Disseminated sulphides: <40% sulphide. Disseminated sulphide mineralisation occurs as either heavily disseminated chalcopyrite or disseminated/blebby sulphides with 0.5cm to 1.0cm diameter sulphide blebs with variable pyrrhotite, chalcopyrite and pentlandite contents.

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Although the ratio between the primary sulphide phases changes between, and within, the sulphiderich and sulphide-poor zones, sulphide mineralisation consists of a consistent assemblage of pyrrhotite-pentlandite-chalcopyrite +/- pyrite. Sulphide content and metal grade are well correlated, with higher sulphide concentration corresponding to higher metal content.

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Figure 6. 3D view (looking NNE) – Gonneville Intrusion domains, sulphide zones and dolerite dykes.

3.3 Mineral Resources

The Resource has been independently prepared by leading mining and geological consultants Cube Consulting. The Resource has been reported in accordance with the 2012 Edition of the Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves (JORC Code), is effective 23 April 2024, and is shown in Table 1.

Table 1. Gonneville Mineral Resource Estimate (Resource) 23 April 2024

Classification	Mass	Grade				Contained	Metal
	Mt	3E (g/t)	Ni (%)	Cu (%)	Co (%)	3E (Moz)	Ni (kt)	Cu (kt)	Co (kt)
Measured	2.9	1.20	0.21	0.17	0.018	0.12	6.1	4.8	0.52
Indicated	400	0.79	0.15	0.087	0.015	10	610	370	65
Inferred	250	0.80	0.15	0.076	0.014	6.4	370	200	37
Total	660	0.79	0.15	0.083	0.015	17	960	540	96

The Resource is reported according to domain (oxide, transitional, fresh) as well as codified confidence levels (Measured, Indicated or Inferred) – refer to ASX Announcement on 23 April 2024 for full details.

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Table 2. Gonneville Mineral Resource Estimate (JORC Code 2012), 23 April 2024.

Domain Cut-off NSR (A$/t)
	Classification Mass	Grade	Contained metal

	(Mt)	Pd (g/t) Pt (g/t) Au (g/t) Ni (%) Cu (%) Co (%)	Pd (Moz) Pt (Moz) Au (Moz) Ni (kt) Cu (kt) Co (kt)
	Measured -	- - - - - -	- - - - - -
	Indicated 7.0	1.9 - 0.05 - - -	0.43 - 0.01 - - -
Oxide – in-pit 25
	Inferred 6.1	0.54 - 0.03 - - -	0.11 - 0.01 - - -

	Subtotal 13	1.3 - 0.04 - - -	0.54 - 0.02 - - -
Sulphide (Transitional) – in-pit 25	Measured 0.4	0.82 0.18 0.03 0.19 0.160 0.020	0.01 0.00 0.00 0.67 0.56 0.07
	Indicated 14	0.68 0.16 0.03 0.16 0.103 0.020	0.30 0.07 0.01 22 14 2.7
	Inferred 0.1	0.72 0.21 0.02 0.13 0.101 0.014	0.00 0.00 0.00 0.19 0.15 0.02
	Subtotal 14	0.69 0.16 0.03 0.16 0.104 0.020	0.32 0.08 0.01 23 15 2.8
	Measured 2.5	1.0 0.22 0.03 0.21 0.168 0.018	0.08 0.02 0.00 5.4 4.3 0.45
	Indicated 380	0.60 0.14 0.02 0.15 0.088 0.015	7.4 1.7 0.30 570 340 57
Sulphide (Fresh) – in-pit 25
	Inferred 240	0.60 0.14 0.02 0.15 0.074 0.015	4.6 1.1 0.15 350 170 35

	Subtotal 620	0.60 0.14 0.02 0.15 0.083 0.015	12 2.8 0.45 930 520 92
Sulphide (Fresh) –	Measured -	- - - - - -	- - - - - -
	Indicated -	- - - - - -	- - - - - -
Mineable Shape 110
	Inferred 7.3	1.7 0.38 0.09 0.16 0.192 0.015	0.40 0.09 0.02 12 14 1.1
Optimiser (MSO)
	Subtotal 7.3	1.7 0.38 0.09 0.16 0.192 0.015	0.40 0.09 0.02 12 14 1.1
	Measured 2.9	0.99 0.21 0.03 0.21 0.167 0.018	0.09 0.02 0.00 6.1 4.8 0.52
	Indicated 400	0.63 0.14 0.02 0.15 0.087 0.015	8.1 1.8 0.32 600 350 60
All
	Inferred 250	0.63 0.14 0.02 0.14 0.076 0.014	5.1 1.1 0.18 360 190 36

	Total 660	0.63 0.14 0.02 0.15 0.083 0.015	13 2.9 0.50 960 540 96

Note some numerical differences may occur due to rounding to 2 significant figures. Includes drill holes drilled up to and including 23 January 2024.

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Figure 7. Plan view of Gonneville Resource block model (sulphide domains) by classification and drilling.

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Figure 8. 3D view (looking north-north-east) of Gonneville Resource block model (sulphide domains) and drilling.

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Table 3. Gonneville Resource grade-tonne table (sulphide domains, excluding oxide), 23 April 2024.

NSR cut-off in pit	NSR cut-of in MSO	Mass		Grade				Contained
A$/t	A$/t	(Mt)	3E (g/t)	Ni (%)	Cu (%)	Co (%)	3E (Moz)	Ni (kt)	Cu (kt)	Co (kt)
15	110	690	0.75	0.15	0.082	0.015	17	1,000	560	100
25	110	640	0.78	0.15	0.085	0.015	16	960	540	96
35	110	530	0.85	0.16	0.092	0.015	15	830	490	82
45	110	390	0.97	0.16	0.11	0.016	12	640	410	63
55	110	270	1.1	0.17	0.12	0.017	9.6	460	330	44
65	110	180	1.3	0.18	0.14	0.017	7.6	330	260	31
75	110	130	1.5	0.19	0.16	0.018	6.1	240	210	23
85	110	95	1.7	0.19	0.18	0.018	5.1	180	170	17
95	110	73	1.8	0.20	0.19	0.019	4.3	150	140	14
105	110	58	2.0	0.20	0.21	0.019	3.7	120	120	11
115	110	47	2.2	0.21	0.22	0.019	3.3	99	110	9.0
125	110	40	2.3	0.21	0.23	0.019	2.9	84	93	7.6
135	110	34	2.4	0.21	0.24	0.019	2.7	74	83	6.6
145	110	30	2.5	0.22	0.25	0.019	2.4	65	75	5.8
155	110	27	2.6	0.22	0.26	0.019	2.2	58	68	5.1

Note: the grade-tonnage table includes material classified as Inferred, where data is insufficient to allow the geological grade and continuity to be confidently interpreted. The gradetonnage curve excludes oxide domains. NSR is calculated at the Resource price deck, which differs to the mine design and PFS price deck.

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4. Geotechnical

Open-pit geotechnical investigations were undertaken by specialist geotechnical consultants Dempers and Seymour (D&S), who have been engaged on the Project since inception, providing services related to data collection, rock mass modelling and stability analysis. D&S has undertaken a robust geotechnical assessment of the pit slopes for the purposes of PFS design, utilising core logging from the Resource drilling database, complemented by dedicated geotechnical drillholes and laboratory testing to inform a 3D mining rock mass block model (MRMM) and significant geotechnical structural model (SGSM).

These models have been used to develop design slope parameters for each geotechnical domain. Stability analysis has been conducted using Limit Equilibrium (LE) and Finite Element Analysis (FEA). Designed slopes meet the Factor of Safety (FoS) requirements defined by the Consultants based on industry best practice[4] .

For the PFS, a larger and deeper open-pit geotechnical review and pit slope design was completed. The Project geotechnical data source is 23,346 metres of rock mass logging from 58 drillholes. From these, the 3D SGSM and the MRMM were updated.

4.1 Data collection

Geotechnical data was collected from these 58 drillholes, from within the Resource. Selected drillholes were such that the maximum spacing between drillholes was ~200m and a majority of the drillholes were within 100m spacing. Logging of drillholes comprised in-field logging and photogrammetry logging. Photogrammetry logging of additional Resource drillholes was also undertaken as part of the PFS. Logging was undertaken using domain logging methodology developed by D&S. Logging of lithology was undertaken internally by Chalice.

Selected samples from each lithology collected during the drilling program were sent to National Association of Testing Authorities (NATA) accredited laboratory SGS Australia for analysis. Testing undertaken included Unconfirmed Compressive Strength (UCS) with modulus, direct shear, particle size distribution and Atterberg Limit. Additional data collection and laboratory testing is planned as part of the upcoming Feasibility Study (FS).

4.2 Rock Mass Strength

Data collection results were interpreted by D&S to determine Rock Mass Rating (RMR) and Modified Rock Mass Rating[5] (MRMR). RMR was determined though assessment of intact rock strength (IRS), drill core Rock Quality Designation (RQD), joint spacing (Js), fracture frequency (FF) per joint set and joint condition (Jc). Conversion of RMR to MRMR accounted for weathering of rock mass, orientation of joint sets, induced stress and blasting effects. Generally, the units within the MRMM can be classified as having:

« A material range of ~35-37 giving a rating of Poor for transitional material
« A range of ~46-59 giving a rating of Fair to Good for rock units
« A range of ~26-28 giving a rating of Poor for geotechnical structures

Weathered unit has been assigned soil strength parameters consistent with expected behaviours. The spatial distribution of MRMR is shown in Figure 9.

4 Read, J. Stacey, P. (2009) Guidelines for Open-Pit Slope Design. CSIRO Publishing, 2012 Edition.

5 Laubscher, D.H. (1990). A geomechanics classification system for the rating of rock mass in mine design. Journal of the South African Institute of Mining and Metallurgy. Vol 90, No 10, pp 257-273 October 1990.

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Intact rock strength was determined from UCS tests on intact core samples with assigned rock strength being:

« Transitional material: 25MPa-50MPa
« Rock units: ranging per lithology from ~50MPa-180MPa
« Geotechnical structures: 5MPa-50MPa consistent with hard rock within soft rockmass

Structural data from orientated core measurements was assessed using stereonet representations to determine set planes per major rock unit. A global stereonet identified 10 major joint sets predominantly trending to the north and plunging to the west.

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Figure 9. Global Stereonet – joint set data

4.3 Output data

The SGSM was developed to incorporate structures with the potential to adversely affect pit slopes in the stability analysis. Lithology model wireframes were provided from Chalice to D&S and the 3D MRMM was created from drillhole logging data and lab testing. SGSM and the MRMM were validated against the drillhole database, and a 3D model of rock bridge and discontinuities was developed to support assignment of Barton shear strength properties, Hoek-Brown failure criteria or Mohr Coulomb parameters to each block model unit for stability analysis. Figure 10 shows the representation.

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----- Start of picture text -----

Domain W1
Domain E1
Domain W2
Domain W3 Domain E2
Domain E3
Domain W4
----- End of picture text -----

Figure 10. Plan view showing MRMR in the pit shell walls and geotechnical domains

4.4 Pit slope design

The PFS slope configuration designs have been developed by D&S and confirmed to meet PFS geotechnical design standards. The configuration is 10m batter height for weathered/transitional material and 20m for rock units which aligns with industry practice. 25m geotechnical berms have been incorporated where inter ramp height exceeds 80m. Batter angle and berm width has been varied per geotechnical domain with the resultant Overall Slope Angles (OSA) to be:

« OSA range from 37° to 47° for weathered/transition material
« OSA range from 48° to 54° for rock units

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These results are consistent with industry expectations for pit design inputs.

A further review of pit slope configuration will be undertaken in the FS following additional data collection and stability analysis.

4.5 Stability analysis

Stability analysis undertaken by D&S against an acceptance criterion defined by industry best practice[6] demonstrated that the selected slope parameters meet the minimum FoS. Geotechnical outputs from the data collected were used to assign geotechnical strength parameters. Slopes have been modelled as depressurised based on the limited groundwater identified onsite from Resource drilling. Assessed slope geometry from the pit shell based on the geotechnical design parameters was provided to D&S. Deterministic analysis was undertaken using FoS from limit equilibrium modelling and stress reduction factor (SRF) - a proxy for FoS - from FEA.

In the FS, a hydrogeological model will be developed to complement the MRMM; additionally further 3D numerical stability analysis will be undertaken.

5. Mining and Ore Reserves

Entech Pty Ltd (Entech) was engaged to provide mining engineering services for the Study, working together with Chalice. The Geology and Mineral Resource input data delivered to Entech were provided by Chalice and technical consulting firm Cube.

The Study was based on conventional bulk open-pit truck-excavator mining methods, comprising drilling, blasting, loading and hauling. It is assumed that all open-pit mining activities are contracted to a suitable mining contractor. Initial pre-strip / ore stockpiling and mobilisation / setup activity is considered pre-production capital. All mining costs post first production are considered operating costs except the cutback required for the Stage 2 processing ramp up which is considered development capital.

The Mineral Resource starts at surface and has a very large footprint, making it ideal for open-pit mining and contains a blend of oxide, transitional and sulphide mineralisation. No unclassified mineralisation outside of the current Resource is included in the modelling. A transition to bulk underground mining in the longer term is considered likely, however it was not considered in the scope of the PFS.

Mining optimisations were run on a normalised, re-blocked Resource model, in which blocks were diluted, converted into quantities of payable concentrates and a Net Smelter Return (NSR) value was assigned to each ore block, according to:

« Dilution,
« Recovery algorithms,
« Payability algorithms,
« Transport and refining charges,
« State royalties, and
« Mine design metal price and foreign exchange rate assumptions (Table 4).

Table 4. Mine design and economic cut-off macro-economic assumptions.

Mine design / cut-off input	Assumption
Pd (US$/oz)	1,050
Ni (US$/t)	16,500

6 Read, J. Stacey, P. (2009) Guidelines for Open Pit Slope Design. CSIRO Publishing, 2012 Edition.

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Mine design / cut-off input	Assumption
Cu (US$/t)	9,000
Pt (US$/oz)	1,000
Au (US$/oz)	2,200
Co (US$/t)	30,000
AUD/USD	0.65

The mine design macro-economic assumptions influence ultimate pit shell, mine design, economic cut-off and mining inventory, and are deliberately more conservative than the long-term average prices used to evaluate financial metrics. This conservatism is applied to reflect the cyclical nature of commodity prices, ensuring profitability of the mine ‘through-the-cycle’.

The mine design prices used reflect Chalice’s estimated ‘trough’ or low-point of future commodity price cycles, in real terms (2025). Historical price trends and assessment of ‘resistance’ points in the industry cost curves (typically 60-80th percentile) were used to guide the mine design prices.

Blocks were categorised for processing, stockpiling or waste according to their NSR value versus total operating cost of extraction and processing.

During operations, economic cut-off and long-term mine plans are expected to be influenced by changes in macro-economic conditions over time. This may result in significantly more material becoming economic to process in future, which is an inherent advantage of a large, low-grade deposit with a steep grade-tonnage curve.

5.1 Dilution analysis

The Resource block model was re-blocked and regularised to account for ore loss and dilution from open-pit mining. Dilution from dolerite dykes (typically sub-vertical orientation), hangingwall/footwall waste and mineralised waste (below cut-off) was accounted for within the regularisation process.

Dilution and ore loss parameters assumed up to 200t excavators and 10m benches. A selective mining unit (SMU) of 5m x 5m x 5m was selected, which achieved the optimal balance of mining efficiency and selectivity.

The deposit has good mineralised zone strike-dip continuity, with wide mineable widths. Internal dilution is minimal with mineralised waste material, whereas dilution to waste is typical at dolerite dyke boundaries, which are typically sub-vertical.

The regularisation resulted in an overall mining dilution at $A25/t NSR of 4% and ore loss of 4% on ore blocks.

5.2 Open-pit mine design

Pit optimisations were run at mine design prices (revenue factor 0.99), a practical mine design was generated around the shells with 614Mt of total material and a mining production target of 280Mt. The final pit dimensions are 1.7km (strike) x 1.0km (width) x 0.45km (depth). The pit shell is artificially constrained in the North to Chalice-owned farmland, inclusive of a buffer (Figure 11).

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Figure 11. Plan view of the modelled life open pit phases.

This final pit was designed in 8 pit phases, guided by optimisation runs, to achieve optimal grade to the process plant and phasing of mining / push-back costs as much as possible (Figure 12).

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Figure 12. NS section view (looking west) of the modelled open-pit phases.

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Figure 13. 3D view (looking WNW) of modelled open-pit phases 1-2.

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Figure 14. 3D view (looking WNW) of modelled open-pit phases 3-8.

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Figure 15. 3D view (looking ENE) of fresh sulphide ore blocks by NSR within the modelled pit.

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Figure 16. NS Section view (looking West) of ore blocks by NSR within the modelled pit.

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Figure 17. EW Section view 1 (looking North) of ore blocks by NSR within the modelled pit.

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Figure 18. EW Section view 2 (looking North) of ore blocks by NSR within the modelled pit.

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5.3 Open-pit mine schedule

The initial process plant throughput rate in Stage 1 is 1 Mtpa oxide ore and 4Mtpa sulphide ore, which increases to a maximum of 14Mtpa of sulphide ore in Stage 2.

The open-pit mine plan assumes conventional open-pit mining using up to four 200t class excavators loading 150t capacity trucks, blasting on a 10m bench height, with a selective mining unit (SMU) of 5m x 5m x 5m.

Mine schedules have been generated which have aimed to optimise the level of stockpiling and rehandling, in order to deliver the highest achievable feed grade to the process plant, trading off mining costs and hence accelerating cashflows as much as possible. The Net Smelter Return calculation for the purposes of mine scheduling and financial modelling reflects base case evaluation prices.

The initial mining rate ramps up to a total material moved (TMM) of 12Mtpa in Stage 1 (year 2), increasing to a steady state of 38Mtpa in Stage 2 (from year 5). The strip ratio (waste:ore) is 1.2 average over the modelled life (Figure 19).

A total of 94% of material in the open pit mining inventory is in the Measured and Indicated Resource classification category. The cumulative proportion of processed Measured and Indicated Resources remains above 90% until Year 20, well after the expected payback period for both Stage1 and the assumed expansions.

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Figure 19. Annual mining schedule by Resource category and strip ratio profile.

In initial years, higher grade feed is processed, while lower grade feed above cut-off is stockpiled, which enhances early cashflows and returns. Stockpiles reach a maximum of 37Mt in Year 12, and are progressively rehandled to the process plant over time.

Oxide inventory is depleted in Year 9, after which time the feed to process plant is sulphide only (Figure 20).

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Figure 20. Process plant feed schedules by ore type and NSR at long-term PFS assumptions

The PFS outlines a large-scale, diversified metals production profile over ~23 years:

« Stage 1 (Years 1 to 4) : ~151koz 3E, 3.2kt Ni, 5.2kt Cu, 0.3kt Co per annum « Stage 2 (Years 5 to 23) : ~238koz 3E, 7.7kt Ni, 8.7kt Cu, 0.7kt Co per annum

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Figure 21. Gonneville 3E precious metal production profile (recovered).

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Figure 22. Gonneville base metal production profile (recovered).

Annualised mining and processing physicals are tabulated in Table 5.

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Table 5. Mining and processing physicals by year. (from commencement of production)

Item	Unit	Yr 0 (CY29)	Yr 1	2	3	4	5	6	7	8	9	10	11	12	13	14	15	16	17	18	19	20	21	22	23
Mining
Oxide ore mined	Mt	4.0	0.3	-	-	2.4	0.3	1.0	-	0.7	0.2													-	-
Sulphide ore mined	Mt	0.1	6.4	7.6	8.2	2.4	9.9	13.1	17.8	10.6	19.0	19.2	18.9	19.5	11.4	7.0	10.5	15.3	16.4	14.8	11.5	8.8	9.6	8.7	4.2
Waste mined	Mt	7.9	5.4	4.4	3.9	17.1	12.8	23.0	20.7	27.0	19.2	19.2	19.5	18.8	20.5	21.7	18.4	13.4	12.3	13.9	10.0	7.8	7.0	5.9	3.4
Processing
Production Stage			1	1	1	1	2	2	2	2	2	2	2	2	2	2	2	2	2	2	2	2	2	2	2
Oxide processed	Mt		0.8	1.0	1.0	1.0	1.0	1.0	1.0	1.0	1.0
« Pd grade	g/t		3.32	1.98	1.47	1.90	1.37	1.42	1.13	1.21	0.99
« Au grade	g/t		0.09	0.07	0.05	0.08	0.05	0.05	0.04	0.04	0.05
Sulphide processed	Mt		3.2	4.0	4.0	4.0	12.0	12.0	12.0	12.0	12.0	14.0	14.0	14.0	14.0	14.0	14.0	14.0	14.0	14.0	14.0	14.0	14.0	14.0	13.5
« Pd grade	g/t		1.07	1.07	1.16	0.68	0.64	0.70	0.70	0.59	0.68	0.70	0.72	0.76	0.66	0.55	0.54	0.61	0.62	0.59	0.58	0.60	0.59	0.58	0.55
« Pt grade	g/t		0.23	0.24	0.26	0.15	0.15	0.17	0.15	0.13	0.15	0.16	0.17	0.18	0.15	0.13	0.12	0.13	0.13	0.13	0.13	0.14	0.13	0.14	0.13
« Au grade	g/t		0.05	0.06	0.07	0.03	0.03	0.04	0.03	0.02	0.03	0.02	0.03	0.04	0.03	0.02	0.02	0.02	0.02	0.01	0.01	0.02	0.03	0.03	0.03
« Cu grade	%		0.18	0.18	0.23	0.12	0.10	0.11	0.12	0.09	0.10	0.10	0.10	0.11	0.09	0.07	0.08	0.10	0.09	0.08	0.07	0.07	0.08	0.09	0.07
« Ni grade	%		0.21	0.20	0.19	0.17	0.16	0.15	0.16	0.15	0.16	0.17	0.17	0.17	0.15	0.14	0.14	0.15	0.15	0.15	0.15	0.15	0.15	0.14	0.14
« Co grade	%		0.02	0.02	0.02	0.02	0.02	0.02	0.02	0.02	0.02	0.02	0.02	0.02	0.01	0.01	0.01	0.02	0.02	0.02	0.02	0.01	0.01	0.01	0.01
Cu conc produced	dmt		20.2	28.1	37.4	18.3	41.0	47.9	53.3	36.0	45.9	51.0	52.7	55.2	47.8	34.3	39.6	49.9	44.6	36.4	32.0	36.0	40.6	48.3	31.2
« Pd recovered	koz		34.2	47.3	51.7	25.9	73.1	81.2	83.4	64.4	80.1	97.8	101.9	107.9	90.1	68.8	67.9	80.6	82.0	77.8	75.0	78.1	77.0	75.3	67.0
« Pt recovered	koz		2.9	4.5	5.4	1.3	4.3	6.4	3.5	2.0	3.6	5.4	7.6	8.7	4.6	2.8	1.6	2.2	2.4	2.2	3.3	3.9	3.4	4.2	4.1
« Au recovered	koz		1.9	3.0	4.0	1.3	4.3	6.2	4.5	2.6	3.9	3.8	4.7	6.9	6.1	3.6	2.7	2.8	2.1	1.0	1.4	2.9	4.3	6.2	4.5
« Cu recovered	kt		4.0	5.6	7.5	3.7	8.2	9.6	10.7	7.2	9.2	10.2	10.5	11.0	9.6	6.9	7.9	10.0	8.9	7.3	6.4	7.2	8.1	9.7	6.2
Ni conc produced	dmt		36.8	47.1	44.0	32.8	92.3	83.9	94.7	84.7	98.2	117.5	126.3	118.4	101.3	84.9	83.6	94.7	99.1	102.6	101.3	97.5	94.1	86.4	75.9
« Pd recovered	koz		26.5	34.7	39.0	14.6	39.0	47.2	47.4	31.1	44.5	56.0	59.7	65.7	48.9	30.2	28.4	39.9	41.0	37.1	34.7	38.0	36.9	35.3	28.9
« Pt recovered	koz		6.4	8.9	10.1	4.7	14.1	16.4	14.3	10.9	14.6	18.2	20.6	22.2	16.2	11.8	11.4	13.9	14.0	12.6	12.5	13.0	12.6	13.5	11.9
« Au recovered	koz		1.2	1.9	2.4	0.9	3.1	4.1	3.1	2.0	2.7	2.7	3.1	4.4	4.0	2.6	2.1	2.2	1.7	1.0	1.2	2.1	3.0	4.0	3.1
« Ni recovered	kt		2.9	3.8	3.5	2.6	7.4	6.7	7.6	6.8	7.9	9.4	10.1	9.5	8.1	6.8	6.7	7.6	7.9	8.2	8.1	7.8	7.5	6.9	6.1

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Item	Unit	Yr 1	2	3	4	5	6	7	8	9	10	11	12	13	14	15	16	17	18	19	20	21	22	23
« Co recovered	kt	0.3	0.3	0.3	0.3	0.7	0.7	0.7	0.7	0.8	0.9	0.9	0.9	0.8	0.7	0.7	0.8	0.8	0.8	0.8	0.7	0.7	0.7	0.6
Dore produced	koz	73.7	70.6	64.5	59.3	101.5	106.0	101.8	93.0	97.8	95.6	98.1	101.5	92.7	80.8	80.2	86.5	87.0	84.7	83.5	85.0	85.3	85.3	78.6
« Pd recovered	koz	70.6	67.1	60.9	56.5	97.5	101.2	98.1	90.5	94.3	93.5	95.5	97.6	89.3	78.8	78.7	85.0	85.9	84.2	82.7	83.4	82.8	81.7	76.0
« Au recovered	koz	3.1	3.4	3.6	2.8	4.0	4.8	3.7	2.6	3.5	2.1	2.6	3.9	3.5	2.0	1.5	1.6	1.2	0.6	0.8	1.6	2.4	3.5	2.6
NSR/tonne processed	A$/t	100.0	97.1	103.1	63.1	57.0	61.7	62.2	49.4	59.1	62.1	65.8	69.3	58.5	44.8	44.6	52.4	52.1	48.8	47.4	49.4	50.1	51.9	45.0

Note: NSR calculation at base case prices.

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5.4 Ore Reserve

The PFS outlines a maiden Proved and Probable Ore Reserve for the Project, limited to the open-pit portion of the Measured and Indicated Resources which have demonstrated economic viability (Table 6).

The Ore Reserve was estimated based on the mining method, mine design, and cost input parameters, applied determined in the PFS. A mining schedule containing Ore Reserves only was evaluated to test economic viability, and this test determined that the Ore Reserve was most sensitive to commodity prices. The Ore Reserve mine plan covers 22 years of mine life.

The maiden Ore Reserve is 260Mt @ 0.86g/t 3E (Pd+Pt+Au), 0.23% Ni, 0.098% Cu, 0.017% Co, containing 7.1Moz 3E, 400kt Ni, 250kt Cu, 43kt Co.

Table 6. Gonneville Ore Reserve Estimate (Reserve).

Classification	Mass				Grade	Grade				Contained Metal	Contained Metal	Contained Metal
	Mt	Pd (g/t)	Pt (g/t)	Au (g/t)	Ni (%)	Cu (%)	Co (%)	Pd (Moz)	Pt (Moz)	Au (Moz)	Ni (kt)	Cu (kt)	Co (kt)
Proved	2.5	1.1	0.23	0.03	0.22	0.18	0.018	0.087	0.018	0.0024	5.4	4.4	0.45
Probable	260	0.67	0.15	0.026	0.16	0.098	0.017	5.6	1.3	0.22	400	250	43
Total	260	0.68	0.15	0.026	0.16	0.098	0.017	5.6	1.3	0.22	400	260	43

The Reserve has been prepared by a Competent Person and reported in accordance with the requirements of the JORC Code (2012). All assumptions underpinning the maiden Ore Reserve are contained within JORC Table 1 Section 4 in Appendix A-3.

6. Metallurgy and processing

The PFS commenced in mid-2023 with the aim of developing a simplified, robust and lower risk flowsheet that was well suited to bottom-of-the-cycle commodity prices. The PFS involved an investment of ~A$15M into metallurgical test work and flowsheet design.

Evaluation work included generating over 100 samples derived from 33 dedicated metallurgical drill holes and conducting over 1,400 metallurgical tests. Metallurgical testing has undergone multiple phases and flowsheet configuration iterations since it commenced for the Scoping Study in 2021. To date, flowsheet development test work for the PFS has involved:

« Comminution (crush, grind) testwork utilising conventional Semi-Autogenous Grinding (SAG)-ball milling, High Pressure Grinding Rolls (HPGRs), Vertical Roller Mills (VRMs) and IsaMill[TM] ;
« Froth flotation (concentration) testwork utilising sequential copper-nickel configurations, with a focus on producing saleable smelter-grade concentrates;
« Leach testwork utilising standard gold industry techniques to recover additional palladium and gold from the flotation tails and oxide material; and
« Magnetic separation testwork on oxide and flotation tails, aiming to remove reactive sulphides and therefore reduce leach reagent consumption in the Resin-in-Pulp (RIP) circuit.

The PFS metallurgical and mineralogical testwork and investigations continued from the work completed for the Scoping Study from 2021 through to 2023. Testwork for the PFS program then continued through to Q3 CY25.

All metallurgical and processing studies were managed by the in-house Chalice team. The in-house Chalice team were supported by two third-party laboratories for testwork, three third-party

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engineering houses for process plant scoping, design and flowsheet development and an additional third-party for mineralogy.

6.1 Metallurgical testwork

Flowsheet development flotation and leach testwork was completed on all sulphide and oxide mine composites across both open and locked cycle testwork phases. The sulphide composites were generated from over 100 samples, derived from 33 dedicated metallurgical drill holes (large diameter PQ core) that were completed in 2023-2024 across the Resource (Table 7).

Comprehensive metallurgical testwork programs were completed over two years, which involved:

« Material characterisation and variability
« Flotation testwork
« Open circuit and locked tests for optimised copper and nickel recovery
« Variability testwork of samples and composites at optimised grind size
« Separation of magnetics (predominantly magnetite) from both oxide and sulphide feed
« Precious metals leach and recovery testwork program
« Leaching of precious metals from oxide and flotation tails (including some blended samples)
« Adsorption of leached product on to resin
« Elution of product from resin
« Recovery circuits to smelt a composite doré
« Tailings and filtration testwork on sulphide ore tailings
« Concentrate quality analysis for copper and nickel

6.1.1 Feed samples

The PFS testwork program was performed on samples derived from 22 dedicated PQ diamond drill met holes, in addition to 11 dedicated PQ met holes from the Scoping Study testwork phase. Metallurgical samples comprised:

« Seven mine composites created for flowsheet development and optimisation work.
« 51 variability samples and five variability composites.

The sulphide composites comprise a mix of high-grade (in the early years) and low-grade feed, to provide a representative spectrum of feed for a long-life bulk open-pit mining operation. The samples have sufficient spatial and grade variability to represent the orebody accurately.

Table 7. Gonneville PFS metallurgical composite details.

Composite	No. of samples	Litho-geochemical Domains	Holes selected	Composite grade
HG2 Yr1-4	9	2 Gabbro, 3 Pyroxenite, 4 High-Cr Ultramafic, 5 Serpentinite (Harzburgite)	JDMET020, JDMET025, JDMET029, JDMET030, JDMET032	1.02g/t Pd, 0.21g/t Pt, 0.02g/t Au, 0.27% Ni, 0.23% Cu, 0.03% Co
		1 Serpentinite (Harzburgite),2	JDMET020, JDMET021,	0.83g/t Pd, 0.14g/t
HG4 Yr1-4	15	Gabbro, 3 Pyroxenite, 4 High-Cr Ultramafic, 5	JDMET022, JDMET025, JDMET027, JDMET029,	Pt, 0.03g/t Au, 0.24% Ni, 0.21% Cu, 0.03%
		Serpentinite (Harzburgite)	JDMET030, JDMET032	Co
		1 Serpentinite (Harzburgite),	JDMET019, JDMET021,	1.09g/t Pd, 0.26g/t
HG2 Yr5+	18	2 Gabbro, 3 Pyroxenite, 4 High-Cr Ultramafic, 5	JDMET022, JDMET025, JDMET027, JDMET031,	Pt, 0.09g/t Au, 0.20% Ni, 0.23% Cu, 0.02%
		Serpentinite (Harzburgite)	JDMET032, JDMET033	Co

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Composite	No. of samples	Litho-geochemical Domains	Holes selected	Composite grade
			JDMET014, JDMET016,
		1 Serpentinite (Harzburgite),2	JDMET019, JDMET021,	0.83g/t Pd, 0.16g/t
HG4 Yr5+	40	Gabbro, 3 Pyroxenite,4 High- Cr Ultramafic, 5 Serpentinite	JDMET022, JDMET023, JDMET024, JDMET025,	Pt, 0.04g/t Au, 0.17% Ni, 0.13% Cu, 0.02%
		(Harzburgite)	JDMET027, JDMET028,	Co
			JDMET031, JDMET033
			JDMET013. JDMET014,	0.55g/t Pd, 0.11g/t
LG S21	17	5 Serpentinite (Harzburgite)	JDMET015, JDMET016, JDMET017, JDMET018,	Pt, 0.01g/t Au, 0.16% Ni, 0.07% Cu, 0.014%
			JDMET020, JDMET023	Co
			JDMET013, JDMET014,	0.58g/t Pd, 0.15g/t
LG CR2 Nov	10	4 High-Cr Ultramafic	JDMET015, JDMET018, JDMET019, JDMET020,	Pt, 0.01g/t Au, 0.17% Ni, 0.10% Cu, 0.02%
			JDMET023, JDMET024	Co
LG PYX C2	13	3 Pyroxenite	JDMET013, JDMET022, JDMET023 JDMET025, JDMET026, JDMET027	0.65g/t Pd, 0.12g/t Pt, 0.05g/t Au, 0.15% Ni, 0.15% Cu, 0.02% Co
			JDMET014, JDMET017,	1.78g/t Pd, 0.56g/t
Oxide MC	7	N/A	JDMET019, JDMET020, JDMET023, JDMET026,	Pt, 0.05g/t Au, 0.17% Ni, 0.23% Cu, 0.08%
			JDMET028	Co
LG nickel	12	1 Serpentinite (Harzburgite),2 Gabbro, 3 Pyroxenite,4 High- Cr Ultramafic,	JDMET20, JDMET021, JDMET022, JDMET025, JDMET027, JDMET029, JDMET030, JDMET0322	0.55g/t Pd, 0.13g/t Pt, 0.03g/t Au, 0.15% Ni, 0.15% Cu, 0.02% Co
				1.75g/t Pd, 0.28g/t
Extreme LG nickel	7	2 Gabbro, 3 Pyroxenite,4 High-Cr Ultramafic,	JDMET019, JDMET025, JDMET029, JDMET032	Pt, 0.06g/t Au, 0.12% Ni, 0.11% Cu, 0.02%
				Co
Extreme LG copper and sulphur	8	1 Serpentinite (Harzburgite)	JDMET021, JDMET022, JDMET027	0.35g/t Pd, 0.09g/t Pt, 0.01g/t Au, 0.16% Ni, 0.02% Cu, 0.02% Co, 0.22% S
		1 Serpentinite (Harzburgite),	JDMET014, JDMET019,	0.47g/t Pd, 0.11g/t
Low sulphur	21	3 Pyroxenite, 4 High-Cr Ultramafic, 6 High-Cr	JDMET021, JDMET022, JDMET025, JDMET027,	Pt, 0.02g/t Au, 0.17% Ni, 0.05% Cu, 0.04%
		Ultramafic	JDMET028, JDMET029	Co, 0.39% S
		3 Pyroxenite, 4 High-Cr	JDMET014, JDMET020,	1.30/t Pd, 0.21g/t Pt,
High Talc	9	Ultramafic, 6 High-Cr	JDMET022, JDMET023,	0.05g/t Au, 0.15% Ni,
		Ultramafic	JDMET025	0.09% Cu, 0.02% Co

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==> picture [469 x 662] intentionally omitted <==

Figure 23. Distribution of metallurgical samples in the open-pit design.

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6.1.2 Process water samples

All testwork to date has been conducted using Perth tap water which typically ranges between 200600 mg/L total dissolved solids (TDS). The PFS assumes process water will be sourced from the Alkimos Waste Water Treatment Plant (WWTP), which has a typical TDS range of 590-670mg/L, which is considered comparable.

6.2 Mineralogy

6.2.1 Sulphide

Mineralogical analyses used over 50 variability samples and were carried out on litho-geochemical and mining composites, flotation concentrates, tailings and a sulphide ore magnetic concentrate. The variability and composite works were carried out with grind to a P80 53μm fineness, using the technically superior, faster and lower cost automated mineralogy solution that uses Scanning Electron Microscopy (TESCAN TIMA), which represents a simpler solution than that used in the Scoping Study work.

Reporting parameters included mineral abundance, elemental (Ni, Cu, S, Fe) deportment, grain and particle size distribution, mineral liberation, mineral association, mineral composition of particles, grade recovery, mineral locking, PGM speciation, PGM mineral association and particle shape (L:D ratio silicates).

Key mineralogical findings included:

« The platinum group metals occur in two main forms:
« In platinum group minerals as interstitial grains on the boundary of sulphides or silicates; and
« In solid solution in the sulphide minerals, predominantly occurring in pentlandite.
« Platinum group minerals host >90% of the Pd and >98% of the Pt:
« 75% of the Pd is in palladium-bismuth-tellurides and 15% is within other palladium minerals.
« >50% of palladium minerals are associated on the boundary of sulphides, whereas >55% of platinum minerals are associated on the boundary of silicates.
« Where the platinum group minerals occur as grains in the interstices of sulphides or silicates, they tend to be very fine (typically P50 of <10µm and P80 of < 20µm) and are liberated during milling.
« Platinum group mineral grain size was consistent across the samples tested, with an average grain size of 4–6 µm, and most grains falling between 3µm and 13µm.
« <10% of the Pd is within solid solution in sulphide minerals, predominantly occurring in pentlandite.
« 10-20% of the nickel is in non-sulphide form, with almost all the silicate-contained nickel present as ultra-fine nickel sulphide minerals within the silicates.
« A proportion of the nickel, particularly in-low grade samples, is not available to conventional recovery. Low nickel grades are therefore likely to be associated with low pentlandite to pyrite ratios, which will lead to lower nickel concentrate grades / recovery for this material.
« The cobalt is associated with the nickel and iron sulphides.
« The dominant copper mineral is chalcopyrite.
« X-ray diffraction (XRD) analysis shows variable amounts of magnetics, ranging from ~1% to ~20%, but averaging ≥12%.
« Serpentine, chlorite and amphiboles are the dominant non-sulphide minerals, with varying amounts of talc.
« Pyrite and pyrrhotite are the predominant sulphide gangue minerals.

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Figure 24. Mineral abundance across the variability samples.

« The copper and nickel are fine grained with a median nickel grain size of 10.8μm and copper grain size of 12.8μm for a P80 53μm primary grind.

Given the large quantity of Pd and Pt in individual mineral species (platinum group minerals) as opposed to within sulphide minerals, a precious metal CN leaching and recovery circuit was investigated to produce a precious-metals doré. This abundance of palladium and platinum in individual minerals rather than in sulphide / refractory form is highly unusual for this style of mineral deposit.

6.2.2 Oxide

Magnetite / hematite, chlorite and clays are the most abundant minerals in the oxide samples.

There were few platinum group mineral grains found in samples, indicating that the minerals present in the oxide zone are finer than in the sulphide ore. This helps explain why the oxide platinum group minerals cannot be recovered with froth flotation.

6.3 Process and flowsheet design

Several processing flowsheets have been investigated, with the aim of maximising metallurgical recoveries and metal payabilities whilst minimising costs and risk. Given the large scale of the Resource and unique characteristics of the Project location, multiple flowsheet iterations and optimisations have been completed since 2021.

Since the August 2023 Scoping Study, Chalice had a major metallurgical breakthrough and redesigned the flowsheet to produce saleable nickel concentrate (>6% Ni) from low-grade samples – something thought unachievable during the Scoping Study testwork phase.

This breakthrough simplified and optimised the process flowsheet for the Project considerably, removing the need for a hydrometallurgical process for the nickel concentrate – which materially reduces execution risk, piloting requirements as well as capital and operating costs.

A summary of Project scope and expected output changes between the 2023 Scoping Study and the new flowsheet are listed below (Table 8):

Table 8. New flowsheet impacts relative to 2023 Scoping Study.

Item	Impact of current flowsheet, relative to the 2023 Scoping Study
Capital	Significant reduction due to removal of hydrometallurgical process.
costs/intensity

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Item	Impact of current flowsheet, relative to the 2023 Scoping Study
Operating costs	Significant reduction in unit operating costs due to removal of the hydrometallurgical
	process and reduction of leach reagent consumption.
	No material change expected for other processes.
Sulphide recoveries	Marginally lower overall recoveries but outweighed by expected reduction in costs –
(indicative)	testwork and optimisation continues, with the potential to further improve recoveries.
Payabilities	Marginally lower Ni-Co payabilities through selling concentrate versus mixed hydroxide
	precipitate (MHP), but superior outcome due to less complex circuits, lower capital
	and expected reduction in operating costs.
Complexity/risk	Materially reduced, utilising simple, proven, industry-standard technology.

Further optimisation work will continue through the FS with trade-off studies evaluating input costs (e.g. grind size, reagent use) versus metal recoveries and payabilities at a range of commodity prices.

The processing plant will produce three saleable products: copper-palladium-platinum-gold and nickel-cobalt-palladium-platinum concentrates, and palladium-platinum-gold doré, utilising industry standard processing equipment.

Individual circuit composition comprises:

« Oxide comminution circuit comprising crushing, scrubbing grinding and classification.
« Sulphide comminution circuit comprising jaw crushing, SAG, Ball and IsaMill[TM] grinding mills and classification. In Stage 2, a Gyratory crusher/SAG and ball mill and tertiary IsaMill[TM] configuration is used.
« Selective Cu-Ni flotation: Conventional sulphide sequential flotation comprising two Cu-cleaning stages to a concentrate grade of 20% Cu and three Ni-cleaning stages to a concentrate grade of 8% Ni for offtake to copper and nickel smelters. Precious metals report to both concentrates and cobalt reports to the nickel concentrate.
« Precious metal leach: Sulphide flotation tails and oxide feed are processed in a cyanide (CN) leach circuit, utilising resins (RIP) to recover product to a precious metal doré. Palladium and gold are recovered from the oxide and sulphide, whilst for the sulphide small amounts of platinum and nickel are also recovered in this circuit.

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Figure 25. Schematic Process Block Flow Diagram (Stage 1).

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6.3.1 Trade-off studies

Several techno-economic trade-off studies were completed as part of the PFS, in order to achieve the optimal value-risk solution trading off recovery, capital costs and operating costs.

Table 9. Process flowsheet trade-off studies

Trade-off scope Scenarios considered Work conducted
	Conclusion

Nickel hydromet circuit to produce MHP vs selling a Ni concentrate Albion / POX/ Activox / Platsol / Kell/ ammonia leach vs Ni flotation only Completed ~800 batch flotation tests and ~29 locked cycle flotation tests. Leach testwork conducted on each of the hydromet scenarios considered methods + a high-level trade-off study. Downstream processing producing MHP delivered sub-economic returns at reserve pricing and current payability levels. Nickel flotation producing a nickel smelter concentrate selected. Hydromet treatment options are incentivised in low capital/operating cost environments and/or where MHP trades at a material premium relative to concentrate to offset the additional cost, risk and complexity. Considered a potential long term upside opportunity.
Target primary grind size 38-106μm Full flowsheet mass balances were conducted on the seven sulphide composites to determine the optimal grind size. 38μm typically outperformed a 53μm grind by delivering: - Higher copper grade - Higher copper, nickel and palladium recoveries Total LOM costs for both cases were considered similar, as lower primary grinding OpEx and CapEx costs for 53μm were offset by higher OpEx and CapEx costs for concentrate regrind mills and CN addition requirements.
	38μm selected.
	Mineralogical investigations
	showed relatively fine-grained
	sulphide and PGMs in the orebody,
	which require fine grinding to
	liberate and selectively float into
	separate concentrates. Palladium
	leach extraction was also found to
	be higher at finer grind sizes.




Comminution equipment configuration SAG + Ball + Tertiary grind vs HPGR vs VRM Comminution testwork was conducted on 74 variability samples and three composites. HPGR and VRM pilot plants were also conducted. Alternative grinding technologies incurred additional CapEx and OpEx costs but did not deliver notable improvements in flotation performance. SAG + Ball + Tertiary grind selected. Reduced risk and lowest capital solution. At the assumed project power prices, alternative grinding technologies were sub-economic.
Trade-off studies highlighted that IsaMillslTMincur a lower installed
Tertiar rindin IsaMillTMvs VTM power cost/CapEx that is only	IsaMillTM selected.
y gg equipment mill vs BGRIMM partially offset by a higher OpEx cost when compared to VTMs. BGRIMM mills were found to be too small for the required duty.	A proven technology with superior
	economics.

Flotation configuration Bulk vs split flotation Bulk Ni-Cu-PGE vs split flotation A bulk flotation concentrate was generated in testwork which could be sold to a nickel smelter. However, the combination of precious metals and recoverable Sequential flotation to separate concentrates selected. Sequential flotation into separate copper and nickel concentrates resultsinsignificantlyimproved

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Trade-off scope Scenarios considered Work conducted
	Conclusion

nickel:copper ratios result in poor offtake terms Overall precious metal payabilities are significantly improved by producing a separate copper-PGM concentrate to be sold to a copper smelter, as well as a separate nickel-cobalt-PGM concentrate to be sold to a nickel smelter. offtake and payability terms relative to a bulk Ni-Cu concentrate, even when accounting for some minor recovery loss by misreporting.
Magnetic No magnetic searation vs 42 samples and composites have been tested comparing leach performance with and without the mag sep+ mag regrind stage. Pyrrhotite flotation on the magnetic stream showed potential to generate a saleable iron concentrate with 65% iron.
	Magnetic separation included in
	Stage 2 flowsheet only.
	The resultant iron concentrate
	byproduct provides another
separation leach p ma se + Removing the magnetic iron	potential revenue stream,
pre-treatment g p pyrrhotite reduces the amount problematic	diversifying project risk. This is
flotation CN soluble species, improving palladium to base metal ratios, increasing leach recovery and reducing leach OpEx. To further simplify the process, work is still ongoing to determine leach conditions that don’t require mag sep for Stage 2.	considered an optional additional
	product to be included in Stage 2
	(an upside to PFS economics).


Oxide comminution circuit to include scrubbing ahead of ball milling Scrubber + Ball mill vs larger ball mill The scrubber delivers a lower OpEx and provides additional optionality around treating scrubber oversize material. The additional optionality is advantageous due to the variable nature of palladium deportment to the coarse fraction. Scrubber and Ball mill selected. Scrubbing offers a lower operating cost and allows for the option to reject barren coarse material.
Treatment of oxide scrubber oversize material Sulphide circuit vs Oxide circuit Flotation testwork determined that oxide scrubber oversize material can be added to the sulphide plant feed and maintain the same flotation recovery, provided it does not exceed 5% of the total feed.
	Sulphide circuit selected.
	No additional CapEx required
	when scrubber oversize is added
	to the sulphide circuit.

Oxide leaching Blended vs unblended oxide leaching When oxide ore is fed on its own, testwork exhibited poor rheology characteristics that required a more dilute leach. The dilute leach conditions require ~4 times the CN addition to catalyse the palladium leach reaction and achieve a similar result to blended leach tests. Blended leaching selected. Blending the oxide ore was found to be significantly more cost effective due to more favourable rheology in the leach feed requiring lower CN consumption.
Acid Current data suggests ore sorting is not viable but it will be investigated
	No pretreatment selected.
Pre-treatment conditioning vs further during the next phase of	The fine disseminated nature of
prior to flotation Coarse ore flotation vs ore work. Coarse ore flotation pilots were run	the target minerals has resulted in
	the pre-treatment methods tested
sorting vs none in multiple configurations however thefine grainsize of thetarget	to date being unsuccessful.

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Trade-off scope Scenarios considered Work conducted
	Conclusion

minerals limited the success of the treatment. While acidification prior to flotation was found to increase recovery, it made cleaning stages more challenging and increased leaching costs.
Alternative treatment methods Bio heap leach vs flotation Low sulphur levels in the feed and fine-grained target minerals yielded low base metal recovery using bio heap leach. Flotation selected. Bio heap leach sighter testwork was found to be unsuccessful.
Leach configuration CIP vs RIP Loading isotherms and all major process steps of the elution, EW and regeneration process has b ttd i thti lh	Resin in pulp.
	The maximum loading of
	palladium on carbon was found
	too low to commercialise. Pd
een ese usng synec eac liquors.	loadings were roughly 4.5 times
	greater on resin than carbon.

6.3.2 Oxide comminution

The ball mill work index of the oxide ore is softer than the sulphide ore, with a ball mill work index of 8.3 kWh/t.

In-situ size by assays were conducted across the oxide variability samples (Figure 26) to determine the viability of screening out the coarse material and upgrading the palladium grade feeding the plant. While this appears to work well on some samples, the recovery was found to be too low on others. Scrubbing testwork was conducted during the Scoping Study, confirming the ability to scrub the ore.

The results show that the oxide ore can be beneficiated consistently using a scrubber. It has also identified the benefit of scrubbing and assaying oversize material before addition to the sulphide ore crushing circuit. Roughly 5% of the oxide ore can report to the flotation circuit, with no significant detriment to flotation performance.

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Figure 26. Oxide ore in-situ size by assay- Pd.

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6.3.3 Sulphide comminution

PFS comminution testwork has expanded on the detailed comminution testing undertaken on the 18 composite samples during the Scoping Study. The results confirm that the competent and hard sulphide host rocks are serpentinised, resulting in lower wear due to less abrasive characteristics.

The samples tested cover the Reserve pit shell:

« JKMRC[7] Drop Weight (3 composites)
« SMC[8] drop weight tests (3 composites and 74 variability samples)
« Bond ball mill work indices (3 composites and 65 variability samples)
« Bond abrasion work indices (3 composites and 65 variability samples)
« Sighter HPGR pilot testing (1 composite)
« Sighter roller mill pilot testing (1 composite)

The results are summarised in Table 10 below:

Table 10. Sulphide Comminution Data

	DWi kWh/m3	*Ab**	sg	*SCSE kWh/t**	BBWi (kWh/t)	Bond Abrasion Index
Average	11.8	24.0	2.8	13.1	26.5	0.04
Median	11.7	24.0	2.8	13.0	27.4	0.04
StDev	1.4	2.9	0.1	0.9	3.5	0.02
Max	15.0	36.0	2.9	15.1	32.7	0.10
Min	7.8	18.9	2.7	10.6	17.5	0.01
85th percentile	13.1	21.0	2.9	14.0	30.1	0.06
15th percentile	10.4	26.8	2.7	12.3	21.7	0.02

A comminution trade-off study was completed, comparing a conventional SAG, ball milling circuit against HPGR and Roller mill technologies (Table 9). The comminution flowsheet selected for the PFS was one stage of crushing followed by three stages of milling (SAG mill, Ball mill and IsaMills[TM] ) to target a primary grind of P80 38µm. The transfer size to the tertiary grinding circuit is P80 150µm.

Palladium and base metal mineralogy is fine grained, resulting in a P80 38μm primary grind size selection to optimise product grade and recovery.

6.3.4 Sulphide flotation concentration

During the PFS program, ~800 batch flotation tests and ~29 locked cycle flotation tests facilitated flotation flowsheets and recovery algorithms for use in the mine planning process. The flowsheet was developed from seven separate composites, representing a range of mineralisation and grades. As part of the program, flotation regimes were assessed, varying flowsheet configurations, reagents and grind sizes.

Separate flotation of copper and nickel was achieved for each composite and formed the basis of the flotation flowsheet. The reagent scheme consisted of carboxymethyl cellulose (CMC) for silicate gangue depression, Triethylenetramine (TETA) /sulphite for iron sulphide depression (when required) and Sodium Ethyl Xanthate (SEX)/thionocarbamate for sulphide collection.

Recovery of copper to high-grade concentrates (25-30% Cu) was achieved, despite the target copper concentrate grade being 20% Cu.

7 Julius Kruttschnitt Mineral Research Centre

8 SMC Test® is a laboratory comminution test which provides a range of information on the breakage characteristics of rock samples for use in the mining/minerals processing industry.

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A final nickel concentrate grade of 8% Ni was targeted. Some samples identified activated iron and nickel sulphides, requiring depression during copper rougher and cleaner stages.

Typically, two stages of cleaning were required for copper and three stages for nickel.

Platinum group metals and gold are distributed between both the copper and nickel concentrates and delivered excellent recoveries to saleable smelter-grade concentrates across all composites at an optimal primary grind size of P80 38µm (Table 11 and Table 12).

Table 11. Locked cycle flotation copper concentrates produced and recoveries by composite.

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|||||||||||
|---|---|---|---|---|---|---|---|---|---|
|Sulphide|Mass pull|Cu grade|Cu rec.|Pd grade|Pd rec.|Pt grade|Pt rec.|Au grade|Au rec.|
|Composite|(%)|(%)|(%)|(g/t)|(%)|(g/t)|(%)|(g/t)|(%)|
|HG2 Yr1-4|0.66|27.6|84.3|58.3|35.6|2.56|8.40|1.49|42.0|
|HG4 Yr1-4|0.85|19.8|82.3|37.5|36.0|1.49|7.47|1.49|35.3|
|HG2 Yr5+|0.69|25.5|76.7|67.4|43.6|7.41|19.5|7.36|70.0|
|HG4 Yr5+|0.45|22.3|72.7|60.5|30.9|2.43|5.62|4.09|43.4|
|LG S21|0.24|25.3|70.4|51.8|22.2|3.71|7.48|1.75|22.9|
|LG CR2 Nov|0.52|17.0|83.0|39.7|31.0|1.90|5.82|1.03|44.4|
|LG PYX C2*|0.34|25.2|62.9|26.3|14.7|0.83|1.86|4.86|35.7|

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Table 12. Locked cycle flotation nickel concentrates produced and recoveries by composite.

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||||||||||||||
|---|---|---|---|---|---|---|---|---|---|---|---|---|
|Ni|Pd|Pt|Au|Au|Co|Fe :|
|Sulphide|Mass|Ni rec.|Pd rec.|Pt rec.|Co rec.|
|grade|grade|grade|grade|rec.|grade|MgO|
|Composite|pull (%)|(%)|(%)|(g/t)|(%)|(g/t)|(%)|(g/t)|(%)|(%)|(%)|ratio|
|HG2 Yr1-4|1.56|8.82|52.0|14.5|21.0|3.18|24.7|0.25|17.0|0.78|49.4|21.7|
|HG4 Yr1-4|1.47|8.43|54.9|16.0|26.6|4.79|41.6|0.57|23.5|0.75|50.0|9.76|
|HG2 Yr5+|0.96|7.94|40.5|12.6|11.4|5.73|21.1|1.13|15.0|0.78|48.0|2.54|
|HG4 Yr5+|1.04|6.99|43.5|21.4|25.5|7.51|40.5|1.17|28.8|0.70|54.7|4.43|
|LG S21|0.81|7.82|40.2|12.9|19.0|4.18|28.9|0.85|38.2|0.85|39.3|4.63|
|LG CR2 Nov|0.74|7.74|35.9|15.7|17.6|6.56|28.7|0.23|14.1|0.92|40.7|4.43|
|LG PYX C2*|0.68|6.15|26.9|12.8|14.3|5.07|22.6|0.64|9.43|0.69|28.9|2.30|

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Blending of feed is expected in a bulk open- pit mine plan, and hence averaged recoveries and concentrate grades are likely to fall within the ranges stated above. Recoveries to concentrates are expressed as a proportion of mill head grade.* PYX is an open circuit test and typically recoveries and concentrate grade improve under locked cycle conditions.

A further 51 variability samples and six composites were tested under open-cycle conditions to understand Resource variability and stress test the flowsheet. Table 13 and Table 14 outline the stress testing results, to determine flowsheet performance under extreme conditions.

These results indicate that a saleable product can be produced even from very low-grade copper and nickel ore blocks and that problematic samples due to high talc (high talc sample 3.52%, talc average 1.12%) can be managed through blending.

Table 13. Open cycle variability testwork copper concentrates and recoveries by composite.

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||||||||||||
|---|---|---|---|---|---|---|---|---|---|---|
|Target|
|Cu|Cu|Pd|Pd|Pt|Au|Au|
|Mass|rec.|Pt rec.|
|Composite|grade|rec.|grade|rec.|grade|grade|rec.|
|pull (%)|@ 20%|(%)|
|(%)|(%)|(g/t)|(%)|(g/t)|(g/t)|(%)|
|Cu|
|LG Nickel|0.29|27.5|69.9|74.6|38.5|21.1|1.80|4.23|2.70|29.8|
|Extreme LG Nickel|0.30|27.1|71.5|74.7|156.3|29.0|1.73|2.04|6.25|34.8|
|Extreme LG Copper|0.09|12.5|47.4|37.9|70.0|18.1|5.04|5.23|5.49|34.7|
|Low sulphur|0.11|23.2|52.5|N/A|73.9|17.2|4.11|4.09|5.55|30.9|

----- End of picture text -----

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Composite	Mass pull (%)	Cu grade (%)	Cu rec. (%)	Target rec. @ 20% Cu	Pd grade (g/t)	Pd rec. (%)	Pt grade (g/t)	Pt rec. (%)	Au grade (g/t)	Au rec. (%)
High Talc	0.21	22.3	52.4	N/A	173	28.2	1.89	1.86	8.50	33.5

Table 14. Open cycle variability testwork nickel concentrates and recoveries by composite.

Composite	Mass pull (%)	Ni grad e (%)	Ni rec. (%)	Targe t rec. @ 8% Ni	Pd grad e (g/t)	Pd rec. (%)	Pt grad e (g/t)	Pt rec. (%)	Au grad e (g/t)	Au rec. (%)	Co grad e (%)	Co rec. (%)	Fe: MgO Ratio
LG Nickel	0.74	8.40	33.7	34.5	12.8	18.1	4.79	29.1	0.67	19.1	0.92	38.2	5.45
Extreme LG Nickel	0.54	8.00	36.7	28.0	93.6	31.9	17.06	36.9	2.85	29.1	0.93	41.0	3.27
Extreme LG Copper	0.72	9.39	41.3	37.8	12.6	25.9	3.74	30.8	0.55	27.6	0.87	40.3	0.83
Low sulphur	0.60	11.4	39.7	N/A	15.8	20.4	4.46	24.7	0.63	19.4	1.88	30.3	1.98
High Talc	0.43	7.60	20.6	N/A	34.1	11.2	11.7	23.3	0.88	6.98	0.86	22.9	12.0

The variability testwork program is ongoing, with the available results summarised in Table 15. Variability testing on the samples is not subjected to locked cycle testing (LCT). To allow comparison with the algorithm generated recovery (derived from the composites), the average recovery improvements seen from open vs locked circuit results on the flowsheet development composites was added to the open circuit results (+4% for Cu and + 6% for Ni).

Table 15. Variability summary – palladium, copper and nickel performance across the samples

	No samples Head grade (% or g/t) Concentrate grade (% or g/t) Average Recovery (%)
Element
	# Min Max Average Average Target Open Circuit LCT Adj. Modelled

Nickel	51 0.13 1.25 0.20 8.64 8.00 35 41 40
Copper	51 0.01 0.87 0.16 19.9 20.0 60 64 65
Palladium (into Ni &
	51 0.25 7.87 1.04 31.9/ 72.3 20 / 50 41 46 46
Cu concs)

The average results across the 51 samples for copper, nickel and palladium are in line with the recovery algorithm, as shown in Table 15. This result gives confidence that variability in performance (Figure 27) can be effectively managed between concentrate shipments.

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Figure 27. Variability samples open circuit flotation results vs recovery model (open circuit)

6.3.5 Leach

The PFS leach program consisted of ~250 leach tests, plus variability testing. A resin-in-pulp (RIP) configuration was adopted after investigating all potential configurations. Investigations included the effects of:

« CN addition
« Lead nitrate
« Grind size
« Pulp density
« Viscosity modifiers
« Blending with flotation tailings
« Temperature
« pH
« Shear
« Magnetic separation and magnetics re-grind pre-treatment
« Pre-oxidation
« Lime addition vs caustic addition
« Diagnostic leaching

The effect of grind size, temperature, blending and CN concentration were found to be the most significant contributors to palladium recovery. Trade-off studies conducted determined that a fine sulphide primary grind (P80 38μm), together with a leach at 45-50% pulp density to maximise CN concentration, utilising a 4:1 sulphide flotation tails: oxide blend, yielded the most economic minimum blend ratio.

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The flowsheet consists of a bank of three leach reactors that operate with a total retention time of approximately six hours. The leach circuit is followed by the adsorption train, with a bank of nine pulp mixers, with a total residence time of four and half hours.

The palladium adsorption/elution circuit was developed on synthetic liquor representative of end of mine life (low grade) leach solutions. Testwork on the synthetic liquor determined that PGMs adsorbed onto carbon. However, palladium maximum loading capacities were too low and therefore uneconomic.

The resin selected for flowsheet development recovers palladium, gold and platinum as well as nickel and some copper from solution.

==> picture [483 x 234] intentionally omitted <==

Figure 28. Resin contact test

After loading, the resin is subjected to a three-stage elution process before regeneration. Elution Stage 1:

« >98% of the nickel is selectively recovered from the resin.
« Nickel is then precipitated to make a nickel sulphide product grading ~32% Ni before being blended with the nickel flotation concentrate.
« This step recovers an additional ~2.8% of the nickel in the ROM feed.

Elution Stage 2:

« >98% of the copper is selectively removed from the resin.
« Copper rich stream is used for the NaCN detoxification processes.

Elution Stage 3

« Palladium, gold and platinum are eluted and recovered utilising a Zadra-style elution circuit.
« The PGM-Au sludge is then smelted to produce a saleable mixed PGM-Au doré

Resin is regenerated for reuse in the adsorption circuit. The resin is commercially available and marketed for resin-in-pulp duty. Piloting of the leach/elution circuit is planned to commence in 2026 in parallel with the FS.

Across 37 fresh ore variability samples, the average leach residue palladium grade is 0.18g/t, the same as the average residue grade from development mining composites covering the same pit

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shell. The average leach feed NaCN consumption across the dataset is 0.42kg/t of leach feed or roughly 0.38kg/t of ROM feed.

Magnetic separation and magnetics re-grind pre-treatment are included in the Stage 2 flowsheet. Mining composites and variability testwork covering the first 40Mt of plant feed has determined magnetic separation and re-grind provides a marginal recovery improvement, but a more significant improvement for the low grade PYX, CR2 and S21 composites. Variability samples and composites after the initial 40Mt will be tested in the FS to determine ability to remove the magnetic separation flowsheet in Stage 2.

6.4 Geo-met and recovery modelling

Upon completion of flowsheet development, variability testwork and trade-off studies, full mass balance testwork was completed for PFS engineering. The full data-set of PFS results used to generate the recovery algorithms and design the plant can be seen in Appendix A-3.

From this work, metal recovery algorithms (recovery vs head grade) have been developed for oxide, transitional and sulphide blocks. Overall process recoveries by unit process and by stage for each metal are outlined in Table 16.

Table 16. Process flowsheet recoveries by unit process and by stage.

Domain	Metal Unit process recovery (modelled open-pit life avg) Overall recovery
	Flotation (to Cu conc) % Flotation (to Ni conc) % Leach % Stage 1 avg % Stage 2 avg %
Oxide	Palladium - - 50 50 50
	Gold - - 83 83 83
Sulphide	Palladium 30 16 30 83 75
	Platinum 7 24 - 42 30
	Gold 37 25 21 90 82
	Nickel - 38 - 44 38
	Copper 72 - - 77 72
	Cobalt - 37 - 42 37

A significant amount of mineralogical, metallurgical data has been collected to date on the Project to drive geo-metallurgical studies. In the FS, geo-metallurgical studies will focus on determining predictors of recovery and product quality, in order to model recovery for each block as accurately as possible. It is expected that refinements and improvements to recovery algorithms will be made according to mineralogy (particularly gangue mineralogy), structure and alteration.

Determining areas of higher or lower recovery than predicted from grade only will also assist in accurate NSR value modelling and classification of blocks as ore or waste and potentially refine the mine design and mine plan.

6.4.1 Oxide recovery

Oxide ore is processed in a dedicated comminution circuit and then blended with sulphide flotation tails for treatment in the leach circuit. In leach testwork on oxide material, no relationship between recovery and head grade has been established.

Testwork for oxide samples is summarised below.

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Table 17. Oxide palladium head grades and leach recoveries.

	Pd head (g/t)	Pd recovery (%)
JSV001A	1.61	32.0
JSV001B	1.83	53.1
JOV003	2.90	62.5
JOV004	2.39	40.0
JOV006	1.14	70.2
JOV008	1.32	65.0
JOV009	3.61	39.8
Oxide MC	1.89	56.2
Average	2.08	52.4

Table 18. Oxide gold head grades and recoveries.

Sample	Au head (g/t)	Au Recovery (%)
JSV001B	0.05	91.2
JOV003	0.02	86.4
JOV004	0.02	83.1
JOV006	0.02	83.1
JOV008	0.06	64.4
Oxide MC	0.06	88.9
Average	0.04	82.8

Palladium oxide recovery is assumed to be 50% of head grade across all oxide blocks, due to the highly variable nature of palladium recovery seen in the variability samples and no apparent marker dictating recovery. Gold oxide recovery is assumed to be 83% of head grade across all oxide blocks.

6.4.2 Transitional recovery

There is 4.3Mt of transitional feed within the mining inventory, with an average sulphur grade of 0.44%. As such, the impact of transitional material to project economics is considered very low. The transitional material was split into two categories for processing:

« OXM-1: High sulphide transitional material, which is processed through the sulphide circuit (treated as fresh material), but with a degraded flotation recovery assumption.
« OXM-2: Low sulphide transitional material, which represents ~32.6% of the transitional ore at a <0.25% sulphur cut-off grade (averaging ~0.08% sulphur) and is not expected to be recoverable via flotation, hence it is assumed to be treated as oxide material. However, OXM-2 ore is roughly double the ore hardness of the oxide ore, which is too competent to treat in the oxide ball mill. Therefore it is assumed to be processed through the sulphide comminution circuit but bypass the flotation circuit.

The following recoveries were assigned to transitional blocks:

« OXM-1: conservatively modelled to have 50% of the flotation process recovery of equivalent grade fresh sulphide material and operating costs equivalent to fresh sulphide material.
« OXM-2: conservatively modelled to have the same leach performance as the oxide, but have operating costs equivalent to fresh sulphide material.

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6.4.3 Fresh sulphide recovery

Fresh sulphide ore is processed in a dedicated comminution and sequential flotation circuit, after which the flotation tails are then blended with oxide for treatment in the leach circuit. As such, recovery of metals is to all three saleable products.

Flotation recovery data has been adjusted to target a 20% copper concentrate and 8% nickel concentrate, based on commercial trade-offs and marketing engagement with potential offtakers. Recovery and payability algorithms are applied to each diluted block.

Overall palladium recovery (the primary revenue driver of the Project) to all three products was not found to be significantly variable across the composite samples. (Figure 29). LG PYX C2 reflects an open circuit test, which is why the palladium recovery is lower than target.

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Figure 29. Overall palladium recovery vs head grade.

This is an expected result given the unique characteristics of the palladium minerals in the deposit and the nature of the leach circuit acting as a secondary recovery mechanism to the flotation circuit.

Palladium is hosted primarily within individual bismuth-telluride minerals, with the balance hosted within sulphides (predominantly pentlandite). As such, the bismuth-telluride minerals have two opportunities to be recovered, in the copper flotation stage (by entrainment) as well as in the leach circuit (CN soluble minerals). In addition, palladium hosted within pentlandite is recovered in the nickel flotation stage.

Overall nickel-cobalt recovery (a secondary driver) is highly sensitive to head grade and gangue mineralogy. Nickel and cobalt are primarily recovered to the nickel concentrate via flotation, and a minor amount of nickel is recovered in the leach elution circuit, which is then blended into the nickel concentrate. The nickel concentrate typically contains <1% Cu.

Overall copper recovery (a secondary driver) is less sensitive to head grade and gangue mineralogy. Copper is primarily recovered to the copper concentrate via flotation only, and a minor amount copper misreports to the nickel concentrate. The copper concentrate typically contains <1% Ni.

7. Tailings and waste management

TSF Reviews Pty Ltd, a subsidiary of L&MGSPL, completed the PFS Tailing Storage Facility (TSF) design using the following standards and guidelines:

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« Department of Mines and Petroleum (WA), 2013, Tailings storage facilities in Western Australia – code of practice : Resources Safety and Environment Divisions.
« Department of Mines and Petroleum (WA), 2015, Guide to the preparation of a design report for tailings storage facilities : Resources Safety and Environment Divisions.
« Global Tailings Review, 2020. Global Industry Standard on Tailings Management (GISTM) . International Council on Mining and Metals (ICMM), United Nations Environment Programme (UNEP), and Principles for Responsible Investment (PRI).
« ANCOLD, 2019. Guidelines on Tailings Dams – Planning, Design, Construction, Operation and Closure . Australian National Committee on Large Dams, Melbourne, Australia.
« Geoscience Australia, 2018, The 2018 National Seismic Hazard Assessment for Australia: data package, maps and grid values.
« ANCOLD, 2012. Guidelines on the Consequence Categories for Dams . Australian National Committee on Large Dams, Melbourne, Australia.

Data collection for PFS design encompassed:

« A geotechnical site investigation from eight drillholes and 16 test pits excavated from within the impoundment.
« Analysis of selected samples was conducted to assess foundation conditions and suitability for use in embankment construction.
« Hydrogeological investigation comprised 10 boreholes drilled within the TSF footprint with selected holes geotechnically logged.

7.1 Tailings Storage Facility design

The site for the TSF was originally selected as part of the Scoping Study completed in 2021 and confirmed as part of more recent studies of TSF and Tailings Technology Options which were completed in 2022 and 2023. Site investigations were completed in 2024 with PFS reporting in 2025.

The TSF comprises a High-Density Polyethylene (HDPE) and clay double lined valley-storage facility which has been designed to be raised by downstream construction. The TSF is designed for 7 construction stages of which only 6 stages are required for ~280MT of tailings produced from the modelled open-pit phase. The embankment construction materials are predominantly being sourced from within the TSF impoundment which reduces footprint (Figure 30 and Figure 31).

The design meets and in the case of embankment stability and freeboard, exceeds the requirements of the DMPE Code of Practice, GISTM and ANCOLD Guidelines. TSF closure and rehabilitation has been considered in the design and costed.

Lift 1 of the TSF dam is included within the pre-production CapEx estimate, whereas future lifts are included within sustaining CapEx.

Future work in the FS will focus on further site investigation, seismic hazard assessment, undertaking consolidation testing of tailings, surface water management plan and an update of the design to FS level of detail.

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Figure 30. TSF Plan Lift 1A (Year 3).

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Figure 31. TSF Plan Lift 7 (final design lift).

8. Infrastructure and logistics

The PFS assessed a range of process water, power and logistics solutions for the Project, and preferred options were selected based on functionality, cost, executability, and potential social and

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environmental impacts. Given its proximity to the Perth metropolitan area, the Project has a degree of flexibility and optionality in the method and associated cost to connect to infrastructure.

The preferred infrastructure and logistics options scoped and costed for the PFS are:

« Process water supply : a ~63km long water pipeline from the Alkimos (WWTP) to the MDA, along with required pumping stations. Pipeline and pumping sized for Stage 2 process throughput requirements.
« Power/electricity supply : a grid connection to the South-West Interconnected System (SWIS) via a ~27km, 132kV, double-circuit transmission line, along with some generation and storage on the MDA.
« Logistics : heavy vehicle haulage of concentrates ~260km from the MDA to Bunbury Port for Stage 1, and ~125km to the planned Kwinana Bulk Terminal Port for Stage 2. All product shipped to customers in Asia.

All infrastructure will be co-located where possible to reduce the disturbance footprint and impact to local communities. The Project will continue to advance the design, commercial, social, environmental and technical aspects of the Project as part of the FS.

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Figure 32. Project water and power infrastructure layout showing proposed infrastructure routes and alternative alignments.

8.1 Process water supply

Process water is to be supplied by a new water pipeline from the Alkimos WWTP to the MDA. This preferred sourcing option provides the Project with a high-quality and reliable water source. It is one of few unallocated sources of water in the Perth Basin and is not impacted by climate change, as the water volume available is a function of the growing wastewater footprint of the northern metropolitan area.

Water is to be further treated by Ultraviolet (UV) at the WWTP prior to delivery to the MDA. Pump and pipeline sizing has been undertaken according to Stage 2 process throughput peak requirements. A 500mm diameter steel pipeline will be installed along a ~63km route to the MDA.

Elevation gain between the WWTP and the MDA is ~260m and pumping will be undertaken via a single-stage pump station at the boundary of the WWTP. The system minimum design flow rate is 3 GL per year, with maximum design flow rate of 11GL per year.

In Stage 2, an onsite water storage dam will be constructed to balance seasonal variations in water flows from the WWTP while also storing contact water harvested during winter for use in summer.

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A site water balance has been prepared which accounts for water balance with the TSF, contact water harvested onsite and the Project’s use of water in processing and dust suppression. The MDA water balance was simulated, and the results of the analysis indicate an average water requirement of 3GL per year with maximum of 5GL per year in Stage 2.

8.2 Power supply

Electricity will be supplied via a new grid connection from the Muchea Substation, together with onsite solar cells and diesel peaking generation, supported by battery storage. The PFS assessed grid connection only, onsite power generation only or hybrid methods to provide power.

The annual power consumption for Stage 1 is forecast to be ~265GWh, while Stage 2 is ~760GWh. The installed/consumed project power is ~47/35MW for Stage 1 and ~133/95MW for Stage 2. The largest power user is the steadiest load, the comminution circuit.

A specialist power model was used to assess all realistic power supply options, including forecast SIWS market pricing and the 2025 SWIS transmission plan, and rank them based on their technical and economic merits. The preferred options for both Stage 1 and Stage 2 were then studied to determine technical and financial parameters for PFS assumptions.

A hybrid solar-battery, together with grid connection is forecast to be the optimal lowest cost option. A grid connection for Stage 1 was found to be highly advantageous as it allows participation in the Wholesale Electricity Market (WEM).

The key power supply and reticulation inclusions are:

Stage 1 :

« A new double-circuit, 132kV, ~27km long monopole transmission line to feed the Stage 1 Project from the existing Muchea Substation.
« A new modularised, pre-fabricated ~153MW solar system, coupled to a ~629MWh battery storage system, located in the south-eastern part of the MDA.
« ~42MW of diesel peaking generation.

Stage 2:

« Construction of the Clean Energy Link (CEL) Chittering, part of the Phase 2 2025 SWIS Transmission Plan[9] , allows the Project to connect to the planned Terminal T9 – a new terminal in close proximity to the Muchea Substation.
« A stepdown transformer from 330kV to 132kV will be constructed allowing continuity of project transmission voltage.

8.3 Logistics

The PFS considered all reasonable transport routes, with a focus primarily on transporting products offsite for export. Several transport methods were reviewed, including heavy haulage, concentrate pumping and rail, to Bunbury, Geraldton, Esperance or Kwinana Bulk Terminal (KBT) ports.

Approximate annual produced volumes for each of the two concentrates are shown below in Table 19. The modelled life moisture content of both concentrates is 9%.

9 Department of Energy and Economic Development: South West Interconnected System Transmission Plan (September 2025)

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Table 19. Annual concentrate volumes produced (wmt).

Concentrate	Product	Stage ktpa	1	(Years	1	to	4)	Stage ktpa	2	(Years	5	to	23)
Cu-PGM		29						48
Ni-Co-PGM		44						110

The following logistics option was selected for the PFS:

« In Stage 1 concentrates are transported ~260km from the MDA to Bunbury Port (Figure 33) via heavy vehicle haulage. Some local road upgrades are required.
« Products are offloaded into vessels using Berth 5 at Bunbury Port (Figure 34), then shipped to customers in Asia on Handymax sized vessels.
« Stage 1 landside logistics will be provided by an established logistics provider using an established solution presently being used for extractive industries.
« In Stage 2 concentrates are assumed to be trucked and exported via the planned new Kwinana Bulk Terminal Port (~125km from the MDA), loaded into vessels then shipped to customers in Asia on Handymax sized vessels.
« Construction personnel will either commute by shift rotation and be accommodated in temporary accommodation constructed on MDA or do a daily Drive-in Drive-out commute.
« All operations personnel commuting daily to site via bus or light vehicles, with the majority expected to reside locally.

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Figure 33. Port of Bunbury layout map (Source: Southern Ports website).

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Figure 34. Berth 5 bulk materials berth at Bunbury Port (Source: Southern Ports website).

Pumping nickel and copper concentrates to Muchea in a concentrate pipeline was assessed as impractical, due to the low volumes and excessively high pump discharge pressures.

In Stage 1 access to the project from Muchea is limited to General Access Vehicles (GAV). Restricted Access Vehicles (RAV) have been adopted in Stage 2 to reduce truck traffic, logistics costs and increase road safety. Approvals for RAVs will be investigated in the FS.

8.4 Non-process infrastructure (NPI)

Non-process infrastructure to support a mine and process plant has been scoped, designed and costed in the PFS:

« Limited upgrades of local site access roads including surfacing, widening and intersection upgrades.
« Surface water management, including civil works comprising diversions and storage dams.
« Site buildings (offices and ablutions, workshops, warehousing, etc).
« Temporary accommodation for construction.
« Services infrastructure (power, water, air, etc).
« Explosives storage and management facility.
« Fuel storage and distribution.
« Landfill and waste management.

9. Tenure, approvals and stakeholder engagement

9.1 Tenure

Chalice owns the freehold title over the MDA, which extends over ~26km[2] . The MDA is subject to exploration tenure granted under the WA Mining Act 1978 , comprising Exploration Licences E70/5118, E70/5119 and E70/5353.

To progress the Project, it is intended that portions of this exploration tenure coinciding with freehold title areas will ultimately be converted to Mining Lease(s). The area applied for under a Mining Lease will encompass the Gonneville mine footprint and all associated mining and processing facilities and will align with the Mine Development Envelope outlined in the environmental approvals.

The infrastructure corridors for the power, water and concentrate pipelines required to support the mine development will be progressed for approval via a Miscellaneous Licence under the Mining Act 1978 . Several corridor options are currently being investigated, allowing for a degree of flexibility in alignment of this linear infrastructure

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9.2 Environmental approvals

The Project requires approvals under the WA Environmental Protection Act 1986 ( EP Act ) , and the Commonwealth Environment Protection and Biodiversity Conservation Act 1999 ( EPBC Act ) .

Extensive work has been undertaken by Chalice to develop environmental baselines and define the program of environmental surveys and studies required to support formal environmental assessment during the Scoping Study phase of the Project. This continued through the PFS phase of the Project. This program of work has considered key environmental factors such as flora and vegetation, terrestrial and aquatic fauna, and surface and groundwater across the MDA. All surveys meet WA Environmental Protection Authority (EPA) and Commonwealth Government technical guidelines for environmental impact assessment.

The Gonneville Project was formally referred to the EPA in March 2024. The referral defined the Gonneville Project within an approximate 2,240 ha Mine Development Envelope (MDE)[10] , with clearing of no more than 940ha of remnant native vegetation. The Project referral also includes water and power transmission infrastructure to the Gonneville Project, located within an Infrastructure Development Envelope (IDE).

The EPA Chair determined in April 2024 that the proposal requires assessment under the EP Act (Assessment Number APP-002518218). The level of assessment was set as a Public Environmental Review, with a public review period for the Environmental Review Document (ERD) of eight weeks.

The referral includes key project characteristics for the operational and infrastructure elements of the mine development area and their extent, inclusive of open-pit mines and/or underground mine access, waste rock landforms, ore stockpiles, tailings storage, haul roads, ore processing and top soil and subsoil stockpiles and associated non-processing infrastructure such as offices, workshops, temporary accommodation for construction, landfill and waste management, wastewater treatment and storage infrastructure.

Further refinement continues for the infrastructure corridors and mine development envelope elements, and amendments to their extents will be undertaken prior to the submission of the draft assessment documentation.

The Gonneville Project was referred to the Commonwealth in March 2024 and determined to be a controlled action in July 2024. The Project is to be assessed by Public Environment Report (PER) (EPBC 2024/09839).

The EPBC Act seeks to protect Matters of National Environmental Significance (MNES) with the following controlled action provisions being relevant to the Project:

« Listed threatened species and communities (s.18 & s.18A)
« Listed migratory species (s.20 & s20A)
« Commonwealth Land (sections 26 and 27A) (Infrastructure Development Envelope only)

An environmental impact assessment is currently in preparation for the Project. This utilises outputs of the PFS and engineering studies, to inform potential and predicted impacts against the key environmental factors and MNES. During the assessment process and preparation of the ERD and PER, Chalice will apply the mitigation hierarchy of Avoid, Minimise, Rehabilitate and Offset, to help reduce the adverse environmental impacts.

Environmental offsets are routinely applied to proposals subject to WA and Commonwealth environmental impact assessment and approval processes where there is significant residual impact

10 The Mine Development Envelope (MDE) is the same spatial area as the MDA. The MDE is a term used in the environmental assessment process.

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to key environmental values of MNES. Chalice has developed a Biodiversity Strategy for the Gonneville Project that seeks to deliver on the following biodiversity goals:

« To ensure science-based no net loss of species or habitat diversity as a result of any mining operations.
« To strive towards a net positive legacy for significant species and our local community.

Approximately 400ha of Chalice-owned land adjacent to the MDA have been designated as Biodiversity Offset areas. Chalice has developed a detailed implementation plan for the Biodiversity Strategy and is progressing on-the-ground restoration work at these areas (Figure 35 and Figure 36 ).

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Figure 35. Project Mine Development Area and planned offset areas.

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Figure 36. Pilot restoration area establishment

Chalice has partnered with Syrinx, University of New South Wales, Bamford Consulting and Carnaby Crusaders to undertake focussed research on Chuditch and black cockatoo species, the key threatened fauna species whose protection and management is critical to success against the goal of ‘no net loss of species’.

The findings of these studies will inform both the appropriate restoration responses and control management actions for the restoration areas.

The environmental assessment and approvals will continue to be progressed in parallel with the study phases of the Project.

A number of other secondary approvals are required to support the development of the mine, and these will progress during FS from mid-2026. Primary approvals to support a Final Investment Decision (FID) are targeted for H1 CY28. Indicative approval timelines, which govern the overall project development timeline, are estimates only and not all steps in the approvals process are subject to statutory timeframes and could vary to those anticipated.

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Figure 37. Project indicative study and development schedule.

9.3 Native Title, heritage and traditional owner participation

The Project is located within the South West Native Title Settlement area of WA (the “Settlement”). The Settlement resolves native title in the southwest of WA through the establishment of ILUAs between the Noongar People and the WA Government. As a result of the Settlement, the grant of mining tenure for the Project will not be subject to the requirements of the Native Title Act 1993 .

The MDA sits within the Whadjuk ILUA area. As per the cultural heritage management framework under the Settlement, Chalice entered into an Aboriginal heritage agreement with the Whadjuk People Agreement Group in the form of a Noongar Standard Heritage Agreement (NSHA) in 2018. Chalice has been engaging and working with Traditional Owner representatives since mid-2021 to establish strong, collaborative relationships and understand cultural values in the MDA. To date, Chalice has engaged with over 70 Traditional Owners in this work.

In 2023, Chalice issued an Activity Notice to the Whadjuk Aboriginal Corporation and the South West Aboriginal Land and Sea Council (SWALSC) as per NSHA requirements requesting surveys covering all of Chalice’s farmlands in and around the Gonneville Project. These surveys were conducted by Whadjuk Traditional Owners in 2023 and 2024 with support from Dortch Cuthbert (specialist archaeologists and anthropologists).

Cultural heritage places, comprising archaeological material, have been identified within the mine development envelope. No ethnographic sites were identified during the surveys. No further surveys are required in the mine development envelope with the final report identifying no issues to prevent the development of the Project. Further discussions with the Whadjuk will be required to determine if any heritage sites exist along the potential infrastructure corridors and, if so, how to best avoid those sites and/or to minimise any impact to them.

Ongoing communications and relationship building with the Whadjuk is an important component of Project stakeholder management. Chalice will continue to engage with the Whadjuk to strengthen relationships, and increase participation in the Project as it develops, including cultural heritage protection and management activities through the FS and environment assessment process.

9.4 Stakeholder engagement

Since the Gonneville discovery in 2020, Chalice has actively and transparently engaged with local communities to keep people informed about the Project, to build relationships and better understand issues most relevant to the community.

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Stakeholders relevant to the Project are categorised into three groups: local community, traditional owners and government. Within each, there are subgroups with varying levels of interest, influence, and potential impact on the progression and success of the Project. Often within local community and Traditional Owners there is overlap between groups. Specific engagement plans have been developed for each group.

Communication and direct engagement have occurred through direct meetings and briefings, site visits, participation in community events, community newsletters, monthly advertising in local newspapers, social media, Gonneville Project website, and distribution of project information sheets. Chalice also opened an office in Toodyay in 2022 to better service community enquires about the Project through direct engagement with Chalice staff.

Delivery of ‘on-the-ground’ community and stakeholder engagement is led by Chalice’s Senior Community Relations and Communications Advisor, with support from Chalice’s executives, a corporate affairs specialist managing government relations, and an Aboriginal engagement specialist.

To better understand community sentiment around Chalice and a potential future mine, Voconiq was contracted in 2023 to implement a Local Voices survey program over several years. Local Voices is a unique community engagement program developed over 10 years within Australia’s National Science Agency, CSIRO.

The first Anchor survey was launched in March 2023. In 2024, and 2025 shorter ‘pulse’ surveys were conducted to record and track community sentiment. The surveys were advertised online, in ads and via street ‘pop up stalls’ over a six-week period. The results were compiled and were shared with the public.

The 2025 results are broadly consistent with previous years, which gives Chalice confidence in the overall outcomes and the veracity of the survey process. The 2025 survey showed that more than two-thirds of respondents (68%) show moderate to strong support for a future mine development on Chalice-owned farmland, provided that key concerns are addressed.

Results from Local Voices will inform Chalice’s ongoing community engagement and investment programs and will be an important input to social impact assessment and environmental approval processes through the next stages of the Project.

Chalice will continue an increased program of communication and engagement during the next phase of the Project. Chalice understands that there is a need to communicate development, construction and operations options (including multiple alternatives where applicable) as soon as possible, for local community members to stay informed, provide feedback and consider impacts.

9.4.1 Community investment

Chalice has invested more than ~$11M into the local community via local spend and direct contractors. The mine is expected to generate 1,200 full-time-equivalent (FTE) construction jobs and 500 FTE operations jobs and contribute more than $1.5 billion in direct royalties and taxes to the state and federal governments. The workforce will be largely residential, based in and around the surrounding areas of Perth, making it a highly sought after location for workers in the mining industry.

The Company has established a Community Investment Program, which provides sponsorship opportunities to support sporting, education, community and environmental initiatives. These contributions have been carefully considered to make sure the benefits are broad, and results in an immediate return for the local community.

To date, ~$420,000 has been distributed to programs and events in the local region.

Chalice’s focus areas for community investment align with the following key pillars:

« Education – initiatives that advance and improve regional educational opportunities;

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« Environment – initiatives that protect and rehabilitate the environment;
« Community Connection – supporting local opportunities, events and groups to strengthen the community connection within the region.

Although the Project is still in the development phase, Chalice understands that meaningful community support is critical to developing strong community connections and authentic engagement is an essential element to building a social licence to operate.

In addition to existing community investment, Chalice has agreed to provide additional funding to local communities through a community fund once the Project reaches commercial production. The principles behind this fund will be determined in conjunction with local shires, and benefits will be directed to the local communities in the vicinity of the Project. The fund will aim to create lasting benefits for the local community, which are determined in consultation with the community.

10. Development timeline and implementation

Chalice is now progressing the development plan into a Bankable Feasibility Study (FS). The FS will involve optimising the design and undertaking detailed engineering to prepare the Project for a Final Investment Decision (FID) on Stage 1, targeted in H1 CY28.

The Company is targeting submission of the ERDs in mid CY26 using the PFS development plan as the basis for the submission. Importantly, the approval scope will consider the full scale and long-term impacts of the Project, so there is scope to adjust the staging of construction according to macroeconomic conditions.

Offtake negotiations for copper and nickel concentrates will formally commence in CY26, with the aim of securing foundational customers for these products, whilst maintaining flexibility and optionality for as long as possible. Offtake discussions could potentially include linked Project financing, as a favourable source of capital and mechanism for alignment with downstream partners.

An FID is expected to be made in H1 CY28, subject to the finalisation of all key activities:

«	Feasibility Study completed	H1 CY27
«	Offtake agreements executed	H2 CY27
«	Funding sourced	H1 CY28
«	Major environmental approvals	H1 CY28

Following FID, a 1.5 to 2 year engineering, procurement and construction phase is expected, resulting in first production in early 2030 (Figure 38).

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Figure 38. Project overall development schedule.

Chalice intends to continue to build on its internal project capabilities and develop the Project. The Project is forecast to require $820 million in capital investment, including contingency.

The FS will include a full Project Execution Plan (PEP), which will be developed and endorsed by the Chalice Board of Directors. The PEP will provide a detailed plan to ensure the Project is delivered to scope, on time and on budget. The PEP would also outline the commissioning and ramp-up plan to ensure full production is achieved over targeted timeframes.

The PEP will detail the expected individual capital works packages, their scope, interaction with other packages and method for execution. Execution of the PEP will be the responsibility of Chalice’s Chief Operating Officer.

The contracting strategy for the capital works packages is likely to be a combination of:

« Engineering, Procurement and Construction Management (EPCM) – Chalice responsible for overall management, with contractors completing design, procurement, construction under a contract / schedule of rates.
« Engineering, Procurement, Construction (EPC) – i.e. ‘turnkey’ delivery of package, with fixed price, scope, schedule and process guarantees.
« Hybrid EPCM and EPC approach.
« Service / supply / consultancy contract.

The process plant, water pipeline and power supply packages are expected be contracted with an EPC approach, whereas all other packages would be delivered under an EPCM or hybrid model. The implementation strategy would be finalised in the PEP during the FS and as such, are subject to change.

11. Cost estimates

The Project cost estimates have been developed by external specialist consultants (refer to Section 15) using a cost build-up methodology. The estimates are classified as Class 4 estimates in accordance with AACE International Recommended Practice 47R-11.

All estimates are in real 2025 Australian Dollars (AUD) and no escalation has been applied.

11.1 Development capital expenditure (CapEx) estimates

Development CapEx estimates have been generated for both stages of construction – Stage 1 in CY28/29 and Stage 2 in mine year 4 (CY33). The scope and timing of the Stage 2 expansion is flexible, according to prevailing macro-economic conditions in operations.

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All Project costs incurred between Stage 1 FID and first production have been capitalised. Supporting infrastructure costs presented reflect Chalice’s capital contribution, with certain multi-user scopes expected to have government and/or third-party contributions by infrastructure and finance providers.

A contingency allowance of 12.5% has been applied to the direct cost estimates excluding mining, to account for potential uncertainties in scope, schedule and rates at this level of study.

Table 20. Development CapEx estimates summary.

	Stage 1 Stage 2
Type	Description Pre-Production Expansion
	A$M A$M
Mining	Mob, facilities and equipment 8 0
	Stripping pre-prod / expansion cut-back 71 53
	Subtotal mining 79 53
Processing	Comminution 170 340
	Flotation 50 74
	Leaching 86 59
	Magnetic separation 0 56
	Other process infrastructure 36 25
	Subtotal processing 350 550
Tailings	Tailings Storage Facility (Lift 1) 84 0
Infrastructure	Power and water 95 0
Indirect	Mob/demob, temp facilities & services, admin, commissioning, first fills 56 95
	Engineering, Procurement, Construction, Management (EPCM) 74 51
	Owners team 18 19
	Subtotal indirect 150 170
Contingency	12.5% contingency on plant/infrastructure directs 66 69
Total	820 840

Note: all numbers are rounded to two significant figures.

11.2 Sustaining capital expenditure (CapEx) estimates

Sustaining CapEx estimates have been generated for the two stages of operation – Stage 1 in mine years 1-4 and Stage 2 in mine years 5-23. Costs classified as Sustaining CapEx include:

« TSF lifts 2-7 over the modelled open-pit life (but not TSF operating/maintenance costs which are included in OpEx).
« Plant and infrastructure capital equipment replacement / overhauls – this has been calculated on a percentage-of-direct-cost basis and is applied from mine year 7.

Table 21. Sustaining CapEx estimates summary.

Description	Stage 1 (5Mtpa) A$M	Stage 2 (14Mtpa) A$M
Tailings, plant and infrastructure	30	570

Note: all numbers are rounded to two significant figures.

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11.3 Operating expenditure (OpEx) estimates

OpEx estimates have been generated for the two stages of operations – Stage 1 in mine years 1-4 and Stage 2 in mine years 5-23. OpEx estimates are presented and modelled either in terms of average unit rates per tonne moved, tonne of ore processed or wet tonne of concentrate produced (Table 22).

Contractor mining is assumed and therefore an appropriate contractor margin is included. Load and haul costs increase over time as the pit deepens and hence are expressed at the beginning and end of Stage 2. Mining costs incurred after commencement of production are treated as OpEx, with the exception of the expansion cut-back in mine year 4, which is treated as Development CapEx.

Table 22. OpEx estimates summary.

		Stage 1 Stage 2
Type	Description Unit
		(5Mtpa) (14Mtpa)

Mining	Drill and blast A$/t mined 1.30 1.30
	Grade control A$/t mined 0.16 0.15
	Load and haul A$/t mined 3.90 4.10
	Rehandle A$/t mined 0.13 0.30
	Owners G&A A$/t mined 0.14 0.07
	Subtotal mining A$/t mined 4.90-6.30 5.20-7.50
Processing	Comminution A$/t proc 5.30 7.70
	Flotation A$/t proc 3.60 4.20
	Leaching A$/t proc 7.20 3.90
	Magnetic separation A$/t proc - 0.81
	Other process infrastructure A$/t proc 2.50 1.70
	Subtotal processing A$/t proc 19 18
G&A	General and administration A$/t proc 1.70 1.00
Total	Total mine site OpEx A$/t proc 35 32
Logistics	Road and port A$/wt conc 79 37
	Shipping A$/wt conc 38 38

Note: all numbers are rounded to two significant figures.

12. Product marketing and offtake

The Project will produce three saleable products utilising industry standard processing techniques, including:

« Copper-palladium-platinum-gold (Cu-PGM) concentrate;
« Nickel-cobalt-palladium-platinum (Ni-Co-PGM) concentrate; and
« Palladium-platinum-gold doré (PGM doré).

The products are considered industry standard and commercially attractive to a broad range of potential customers. The products are expected to be marketed and sold as follows:

« The Cu-PGM concentrate is expected to be sold directly to copper smelters in Asia and/or Europe, where offtake terms are expected to be highly favourable based on indicative terms received to date. The copper concentrate is expected to have negligible deleterious elements.

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« The Ni-Co-PGM concentrate is expected to be sold directly to nickel smelters or pre-cursor Cathode Active Material (“pCAM”) refineries in Asia, Europe or North America, where offtake terms are expected to be favourable based on indicative terms received to date. The nickel concentrate is expected to have negligible deleterious elements, with a minor penalty for MgO in lower grade later years of the mine plan.
« The PGM doré is expected to be sold directly to a precious metal refinery, where a nominal refining charge will be payable.

It is assumed that the payable metals in the offtake products will be nickel, copper, cobalt, palladium, platinum, and gold, however, the concentrates do contain minor amounts of rhodium, iridium, silver and other minor critical minerals, and the recovery and potential payability of these metals continues to be further investigated.

For the copper and nickel concentrates, several rounds of indicative terms were received from prospective smelters, based on indicative quantities, timing and specification of Gonneville products. The direct engagement with smelters was facilitated by an independent base metal concentrate marketing expert consultant. A summary of the terms received from a wide range of copper and nickel smelters globally are shown in Table 25 and Table 27.

No offtake agreements have been signed for the Project and as such products are 100% uncommitted.

The Project is strongly aligned to western government policy directives and directly addresses the critical minerals dominance of China, Russia, South Africa and Indonesia. As such, there is a strong case for a future effective western premium on products (through either longer-term offtake or higher realised pricing relative to other non-western offtakes).

12.1 Palladium market overview

The primary long-term revenue driver for the Project is expected to be palladium, however nickel and copper are also considered co-products. Palladium, platinum, rhodium, ruthenium, iridium and osmium form a group of elements referred to as the platinum group metals (PGMs).

More than 85% of palladium is used in catalytic converters for petrol engines (predominantly in light vehicles), which convert as much as 90% of the harmful gases in exhaust (hydrocarbons, carbon monoxide and nitrogen dioxide) into less harmful emissions.

As emission standards become stricter, and vehicles increasingly operate in stop/start conditions (i.e. hybrids), palladium loading per vehicle is increasing over time.

Palladium is also used in electronics, dentistry, medicine, hydrogen purification, chemical / catalyst applications and jewellery. Increasingly palladium is used in high performance electronics applications in the defence sector. Palladium is also a key component of fuel cells, which react hydrogen with oxygen to produce electricity and water.

Global demand in 2024 was 10.1Moz, while global mine supply was 9.6Moz (the market was in deficit). The market has remained in deficit conditions for over a decade.

Table 23. Historical palladium supply and demand 2020-2024.

Troy ounces ‘000	2020 2021 2022 2023 2024
Primary supply (mine) 6,196 6,846 5,964 6,597 6,654
Secondary supply (recycled)	3,128 3,339 3,257 2,865 2,940
Total supply 9,324 10,185 9,221 9,462 9,594
Total demand	9,959 10,195 9,913 10,370 10,095
Surplus/(Deficit) (635) (10) (692) (908) (501)

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Source: Johnson Matthey PGM market report May 2025.

12.2 Cu-PGM concentrate

The Cu-PGM concentrate has been confirmed as a commercially attractive product to a range of copper smelters. There are more than 30 copper smelters worldwide that purchase concentrate feed on the open market.

Six copper smelter/refinery complexes with an established PGM refinery (beneficial given the high PGM content within the concentrate) have been identified and engaged by Chalice through marketing investigations in Asia, Europe and North America.

These complexes typically treat PGM-bearing copper concentrates along with secondary materials (such as auto catalysts) and produce a high value PGM product, typically a palladium/platinum sponge or refined metal.

It is expected that this group of specialist copper smelter/refineries will be the most attractive offtakers of the Gonneville concentrate, as they already have the necessary downstream PGM refining capacity in place and off-take arrangements with end-use customers.

Assays of concentrates produced to date indicate a very clean copper smelter concentrate, with negligible levels of deleterious elements. Expected specification ranges for each shipment are provided in Table 24.

Table 24. Cu-PGM concentrate specifications.

Metal	Unit	Expected range
Copper	%	>20%
Nickel	%	0.64-1.0
Palladium	ppm	40-50
Platinum	ppm	2.0-4.5
Rhodium	Ppm	0.05-0.2
Cobalt	%	0.03-0.06
Gold	ppm	1.5-3.4
Silver	ppm	0-10
Sulphur	%	25
Iron	%	25
Arsenic	ppm	<100
Antimony	ppm	40
Bismuth	ppm	<20
Cadmium	ppm	<5
Lead	ppm	<100
Mercury	ppm	0.1
Uranium	ppm	<1
Zinc	ppm	1,000

Indicative terms provided by copper smelters have formed the basis of the offtake assumptions for the PFS. The indicative terms quoted by parties were uniformly high with no penalties envisaged.

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Table 25. Offtake assumptions for each metal in the Cu-PGM concentrate (modelled life avg).

Metal Payability (net of deductions) Treatment Charge	Refining Charge (PGM- Refining Charge
	Au) (Cu)
% US$/dmt conc	US$/oz USc/lb
Palladium 96% 40 Platinum 67% Gold 91% Copper 95%	15
	15
	5
	4

12.3 Ni-Co-PGM concentrate

The Ni-Co-PGM concentrate has been confirmed as a commercially attractive product to a range of nickel smelters. The pre-cursor Cathode Active Material (“pCAM”) refining industry is also considered a potential customer, however limited volumes of nickel sulphide concentrate have been processed to date by these players.

Three nickel smelter/refinery complexes that purchase nickel concentrate feed containing PGMs have been identified and engaged by Chalice through marketing investigations in Asia, Europe and North America.

It is expected that this group of specialist nickel smelter/refineries will be the most attracted to the Gonneville concentrate, as they already have the necessary downstream processing technology and/or offtake arrangements in place with end-product customers.

Assays of concentrates produced to date indicate a very clean nickel smelter concentrate, with negligible levels of deleterious elements and a low MgO content. Expected specification ranges for each shipment are provided in Table 26.

Table 26. Ni-Co-PGM concentrate specifications.

Metal	Unit	Expected range
Nickel	%	>8
Copper	%	0.5-1.2
Palladium	ppm	15-24
Platinum	ppm	4.4-6.3
Cobalt	%	0.76-0.84
Gold	ppm	0.45-1.5
Silver	ppm	0-5
Sulphur	%	27-42
Iron	%	27-41
MgO	%	1.9-12
Arsenic	ppm	<100
Antimony	ppm	1
Bismuth	ppm	<20
Chlorine	ppm	<100
Fluorine	ppm	Below DL
Lead	ppm	<100
Mercury	ppm	<0.1

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Metal	Unit	Expected range
Zinc	ppm	200

Indicative terms provided by nickel smelters have formed the basis of the offtake assumptions for the PFS. The indicative terms quoted by parties were variable, with only a very minor penalty envisaged for MgO in the lower grade, later years of the mine plan.

Table 27. Offtake assumptions for each metal in the Ni-Co-PGM concentrate (modelled life avg).

Metal	Payability (net of deductions)
	%
Palladium	76%
Platinum	64%
Gold	33%
Nickel	80%
Cobalt	55%

There is potential for more attractive offtake terms from the pCAM industry, given the amount of new processing facilities and very high acid prices, which could provide improved offtake terms relative to existing nickel smelters. This will be further investigated within the FS.

The Study PFS has not considered further processing to a high purity nickel hydroxide or sulphate product. However, given shifting market dynamics, further evaluation is warranted over the longer term as commodity markets evolve.

12.4 PGM doré

The PGM doré is a standard precious metal product which can be refined at various precious metal refineries globally. The doré bar will be sent to an LBMA accredited refinery in Australia, Asia or Europe. Indicative terms have formed the basis of the offtake assumptions for the PFS.

Table 28. Offtake assumptions for each metal in the PGM doré (modelled life avg).

Metal	Payability	Refining Charge
	%	US$/oz
Palladium	99	15
Gold	99	5

13. Financial analysis

A detailed Project Financial Model has been developed to support the evaluation of the Project. The model is purpose-built and includes only unleveraged cash flows directly attributable to the Project, modelled on a 100% basis. The model does not incorporate any assumptions related to the financing structure of the Project.

The financial model has been developed with key financial metrics expressed in real terms over the Project’s 23-year open-pit life. Net Present Values (NPVs) are calculated at FID.

13.1 Key assumptions

Long term commodity price assumptions used in the PFS have been derived from a combination of sources, including industry cost curves, long term consensus forecasts from banks and economic forecasting houses, LME and NYMEX metals futures markets and other industry benchmarks.

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Base case price assumptions are considered conservative and realistic in the current macroeconomic environment, noting that all long term price assumptions are at or below the approximated 95[th] percentile of industry cost curves[11] . Study financial outcomes at spot commodity prices and exchange rate are also presented for reference.

The Australian Government’s Critical Minerals Production Tax Incentive (CMPTI) provides a 10% refundable tax offset on eligible processing and refining costs. The financial model incorporates the CMPTI for 10 years from 2030 for all eligible processing activities.

Table 29 outlines the key assumptions made in the financial analysis of the Project.

Table 29. Key assumptions in Project financial model

Key assumption	Unit	Stage 1 (5Mtpa)	Stage 2 (13-14Mtpa)
Commodity prices (real terms, flat)12
Ni	US$/t	18,750
Cu	US$/t	10,500
Co	US$/t	39,000
Pd	US$/oz		1,300
Pt	US$/oz		1,300
Au	US$/oz		2,900
Financial
WACC (real)	%		8.0
Exchange rate	A$/US$		0.65
Offtake terms (avg)
Copper concentrate
Cu payability	% LME		95
Pd payability	% LBMA		96
Pt payability	% LBMA		67
Au payability	% LBMA		91
Treatment charge	US$/dmt conc		40
Cu refining charge	US$/t Cu		88
Pd/Pt refining charge	US$/oz		15
Au refining charge	US$/oz		5
Nickel concentrate
Ni payability	% LME		80
Co payability	% LME		55
Pd payability	% LBMA		76
Pt payability	% LBMA		64
Au payability	% LBMA		33
PGM doré

11 Wood Mackenzie 2025 nickel and copper cost curves sourced 31 Oct 2025, 95th percentile of palladium cost curve is Sibanye Stillwater US PGM Operations (2025 AISC guidance US$1,320/oz 2E incl S45X credit) sourced 7 Nov 2025).

12 Commodity prices are rounded to two significant figures and are used for the purposes of financial modelling. Commodity prices used in the open pit mine design, optimisation and economic cut-off are different and can be found in the Mining section.

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Key assumption	Unit	Stage 1 (5Mtpa)	Stage 2 (13-14Mtpa)	Stage 2 (13-14Mtpa)	Stage 2 (13-14Mtpa)
Pd-Au payability	% LBMA		99
Pd refining charge	US$/oz		15
Au refining charge	US$/oz		5
Development CapEx estimates
Mining	A$M	79			53
Process Plant	A$M	350			550
Tailings Storage Facility (TSF)	A$M	84			-
Infrastructure	A$M	95			-
Indirect/EPCM/Owners	A$M	150			170
Contingency (12.5% direct)	A$M	66			69
Total Development CapEx	A$M	820			840
Total Sustaining CapEx	A$M	30			570
OpEx estimates (avg)
Mining	A$/t proc	15.00			12.00
Comminution	A$/t proc	5.30			7.70
Flotation	A$/t proc	3.60			4.20
Leaching	A$/t proc	7.20			3.90
Magnetic separation	A$/t proc	-			0.81
Other process infrastructure	A$/t proc	2.50			1.70
General and administration	A$/t proc	1.70			1.00
Total mine site OpEx	A$/t proc	35			32
Logistics (site to smelter)	A$/t proc	1.80			0.85
Taxation
Ni-Cu-Co-Pd-Pt-Au WA Govt royalty rate	%		2.5
Corporate tax rate	%		30
Production Tax Credit	%		10
Schedule
FID	date	Early 2028
Commence Operations	date		2030
Plant ramp up	% throughput	80 yr 1, 100		yr	2+

Note all figures are rounded to two significant figures.

London Metal Exchange
London Bullion Market Association

13.2 Financial return metrics

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Table 30. Gonneville Project Pre-Feasibility Study key financial metrics (open-pit phase only).

Key metric Unit	Base case13	Spot case14
Modelled open-pit life Years 23 >23
Cumulative gross revenue A$bn 16.7 18.1
Cumulative EBITDA A$bn 6.9 8.5
EBITDA margin % 44 49
Cumulative free cashflow (pre-tax) A$bn 4.7 6.2
Cumulative free cashflow (post-tax) A$bn 3.6 4.7
Annual operating cashflow (pre-tax) A$Mpa 280 340
Annual operating cashflow (post-tax) A$Mpa 230 270
NPV8%(pre-tax) A$bn 1.4 2.0
NPV8%(post-tax) A$bn 1.0 1.5
IRR (pre-tax) % 23 29
IRR (post-tax) % 21 26
NPC8%development CapEx A$bn 1.3
Stage 1 payback (from 1stproduction) Years 2.7 2.4
Stage 2 payback (from Yr5) Years 2.5 2.0
All-in Sustaining Costs (AISC)15 US$/oz 3E 370 390

13 Wood Mackenzie 2025 nickel and copper cost curves sourced 31 Oct 2025, 95th percentile of palladium cost curve is Sibanye Stillwater US PGM Operations (2025 AISC guidance US$1,320/oz 2E incl S45X credit) sourced 7 Nov 2025). 14 Spot prices Pd: US$1,500/oz, Pt: US$1,660/oz, Au: US$4,250/oz, Ni: US$14,900/t, Cu: US$12,050/t, Co: US$49,500/t, Cu conc TCRCs US$-40/t, US-4c/lb, Ni conc Ni payability 76%, sourced COMEX, LME, S&P Global 5 Dec 2025.

15 AISC per produced 3E ounce (Pd+Pt+Au), net of byproduct credits after payabilities from Ni, Cu, Co. AISC calculation aligned to the SFA Oxford methodology, which excludes royalties, to compare with PGM industry peers.

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Figure 39. Cashflow profile over modelled open-pit phase (pre-tax, real).

13.3 Detailed metrics by stage

Detailed financial, cost and physical metrics are presented for Stage 1 (the first 4 years), as well as the entire 23 year modelled open-pit phase of the Project, at base case prices (Table 31). Stage 1 has considerably higher feed grade than future years and hence considerably higher margins. All figures are in real terms (2025 AUD) and are unleveraged.

Table 31. Gonneville Project Pre-Feasibility Study detailed metrics by stage (base case prices).

Metric	Unit	Stage 1 (Years 1 to 4)	Modelled open-pit life (Years 1 to 23)
Financial
Gross Revenue (avg)	A$Mpa	450	730
Net Smelter Return per tonne processed	A$/t	90	57
EBITDA (avg)	A$Mpa	260	300
EBITDA margin (avg)	%	61	44
Annual operating cashflow (pre-tax)	A$Mpa	260	280
Annual operating cashflow (post-tax)	A$Mpa	250	230
Capital Costs
Pre-Prod development CapEx (incl. contingency)	A$M	820
Stage 2 expansion CapEx (incl. contingency)	A$M	-	840
Sustaining CapEx	A$Mpa	7	26
Operating Costs (avg)
Mine site cash costs per tonne processed	A$/t	35	32
Mine site cash costs per 3E ounce produced	US$/oz 3E	720	1,130
+ Transport & Selling costs	US$/oz 3E	57	53
- By-product credits (Ni, Cu, Co, Fe)	US$/oz 3E	700	890

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Metric	Unit	Stage 1 (Years 1 to 4)	Modelled open-pit life (Years 1 to 23)
= Total cash costs per 3E ounce	US$/oz 3E	75	290
+ Sustaining costs	US$/oz 3E	32	76
= All-in Sustaining Costs (AISC) per 3E ounce	US$/oz 3E	110	370
PGM Industry Cost Curve Position (net of by-products)	quartile	1st	2nd
Mining Physicals
Total ore mined (excl pre-prod mining)	Mt	27	280
Total waste mined (excl pre-prod mining)	Mt	31	330
Total material movement incl. reclaim (avg)	Mtpa	14	36
Strip ratio (avg)	x	1.1	1.2
Processing Physicals
Total mass processed	Mt	19	280
« Measured	%	9	0
« Indicated	%	91	94
« Inferred	%	0	6
3E (Pd+Pt+Au) grade (avg)	g/t	1.44	0.85
Nickel grade (avg)	%	0.15	0.15
Copper grade (avg)	%	0.14	0.092
Cobalt grade (avg)	%	0.014	0.015
Oxide processing throughput	Mtpa	1	1→0
Oxide modelled life	Years		9
Sulphide processing throughput	Mtpa	4	4→12→14
Sulphide modelled life	Years		23
Produced 3E (Pd+Pt+Au)	koz	600	5,100
Produced nickel	kt	13	160
Produced copper	kt	21	186
Produced cobalt	kt	1.1	15
Pd recovery (avg)	%	71	74
Pt recovery (avg)	%	42	31
Au recovery (avg)	%	88	83
Ni recovery (avg)	%	44	38
Cu recovery (avg)	%	77	72
Co recovery (avg)	%	42	37

Note: values are rounded to 2 significant figures. Gross Revenue is net of payabilities (as invoiced by offtakers)

13.4 Cost profile

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The AISC is intended to highlight the costs and margins of the operation per produced 3E ounce. AISC is calculated as:

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Figure 40. Gross revenue split by commodity and product (after payabilities), avg.

The AISC progressively trends up to US$370/oz 3E over the modelled open-pit life, primarily due to lower overall feed grades over time and higher mining costs as the open-pit gets deeper (Figure 41).

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Figure 41. Gonneville AISC cost vs Gonneville 3E basket price 6F [16] over modelled open-pit phase.

16 Gonneville 3E basket price the weighted average Pd, Pt, Au price after payabilities.

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13.5 Industry competitiveness

The competitiveness of the Project has been assessed against PGM industry peers, whose revenues are driven primarily by platinum or palladium. These mines typically report their cash and sustaining costs per 4E ounce of palladium, platinum, gold and rhodium produced (4E=Pd+Pt+Au+Rh). Byproduct credits from nickel, copper, chrome, cobalt, iridium, ruthenium and other minor metals are offset against costs.

Gonneville’s attractive position on the cost curve highlights a robust and competitive asset that is modelled to be highly profitable through the commodity cycle. The next best peer in the industry has AISC of ~US$721/oz (Impala Canada Lac Des Iles operation in 2024 8F[17] ), over double the predicted Gonneville AISC and has since announced closure plans in mid 2027 (costs artificially low at the end of the mine plan).

17 SFA Oxford 2024 actual, sourced on 4 June 2025

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Figure 42. 2024 PGM industry all-in sustaining cost curve (net of byproduct credits) and Gonneville positioning 7F [18] .

18 Source: 2024 SFA (Oxford) Ltd actual collated costs and revenues used for 4E cost curve data in June 2025. The Gonneville AISC assumes average by-product prices of: Copper US$10,500/t, Nickel US$18,750/t, Co US$39,000/t. AISC calculation aligned to the SFA Oxford methodology which excludes royalties.

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13.6 Sensitivity analysis

« Adapting the mine design / mine plan due to changes in economic cut-off (increasing or decreasing the feed grade to plant and/or overall mine inventory),
« Adapting the process plant to chase higher recoveries through higher reagent use and higher operating costs,
« Applying hedging strategies,
« Increasing plant throughput capacity or performing retrofit / adaptations to the process plant configuration.

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Figure 43. Modelled open-pit phase pre-tax NPV 8 sensitivity analysis.

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Over the mid 2023 to mid 2025 period, the palladium price remained well below the marginal cost of production (~US$1,320/oz) 9F[19] and has only recently recovered to more sustainable levels.

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Figure 44. LBMA palladium price (blue), PFS mine design and PFS long term price, US$/oz.

19 Sibanye Stillwater 2025 guidance for US PGM Operations (US$1,320/oz 2E AISC incl S45X credit), sourced 7 Nov 2025)

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Table 32. Pre-tax NPV and IRR sensitivity to long term Pd, Ni and Cu prices (real).

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« The robustness of palladium demand, particularly from internal combustion and hybrid vehicles and slowing adoption of battery electric vehicles, but also from growing applications in data centres (electronic components, semi-conductors (multilayer ceramic capacitors) and precious metal investment given its extreme scarcity;
« Increasing palladium loadings per vehicle over time as emissions standards become stricter, particularly in the developing world;
« The lack of substitutes, or at least readily available substitutes in palladium applications;
« Palladium demand is extremely inelastic – i.e. consumers are not sensitive to the price, in particular when considering the input cost of palladium in an average internal combustion or hybrid vehicle is currently US$100-200 per vehicle;
« A prolonged subdued price environment for the metal will incentivise new applications (e.g. replacing gold in electrical connector plating, hydrogen production and purification and new chemical / catalytic applications), thus increasing demand over time;
« The instability and challenging investment landscape of Russia, South Africa and Zimbabwe;
« The rapid rise in industry costs in South Africa (>10% p.a.) driven by ageing and deep underground mines, which puts upwards pressure on prices over the long term;
« Structural infrastructure issues, corruption, political instability and high levels of inflation in South Africa and Zimbabwe;

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« Lack of palladium deposits and economically viable development projects globally (supply is extremely inelastic); and,
« Lack of investment in recycled / secondary supply, particularly without a significant, sustained price incentive above current levels.
« Negligible investment was made into recycling or any form of new supply in the last period of sustained incentive prices in 2019-2023, which demonstrates the extreme inelasticity of supply.

13.7 Project funding

Funding is expected to be sourced from range of partners including:

« Western Australian State Government sponsored initiatives
« Australian Federal Government sponsored initiatives
« International government sponsored initiatives
« Offtake partners
« Specialist ‘green’ finance providers
« Commercial banks

« Australian and international governments, from Organisation for Economic Co-operation and Development (OECD) countries, have a strong appetite to support large scale critical minerals projects.
« Export Credit Agencies (ECAs) from major OECD nations, including Australia, Canada, Germany, Japan, and the United States, are showing increasing interest in financing projects such as Gonneville as part of broader national strategies to enhance supply chain resilience.
« ECAs can provide long-term, low-cost debt which bolsters project viability and profile, often catalysing equity participation from sovereign wealth funds, development finance institutions, and strategic investors seeking de-risked exposure to critical minerals.
« The signing of a non-binding Memorandum of Understanding (MoU) with Mitsubishi Corporation in 2024 highlights clear potential for strategic partnership in the development of the Project. This collaboration demonstrates growing global interest in securing reliable sources of critical minerals and reinforces Gonneville’s appeal to world-class industrial counterparties.
« Chalice has a current market capitalisation of approximately A$680 million (at 5 December 2025) and no debt.
« The Company has a strong track record of successfully raising equity funds in a prudent and disciplined manner when required to further the exploration and development of the Project.
« The Chalice Board and management team has experience in mine development, financing and operations in the resources industry.
« Chalice owns 100% of the Gonneville Project and there are no historical financing mechanisms (e.g. royalty, stream, etc) encumbering its development.

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« The Project is located in Western Australia, which is considered one of the lowest-risk, most stable and attractive jurisdictions globally for mining.
« The Project has an initial modelled open-pit life of 23 years.
« Project economic viability has been established at bottom of the cycle commodity prices, with the PFS demonstrating a highly competitive 2[nd] quartile position on the PGM industry cost curve and an unleveraged payback period of under three years – there is considerable financial capacity to cover long term debt repayments.

14. Upside opportunities and risks

14.1 Additional iron byproduct upside

A saleable iron concentrate byproduct was investigated during the PFS, which was inadvertently created in the process of magnetic separation testwork upstream of the leach circuit. The iron concentrate produced had a grade of 65% Fe.

14.1.1 Testwork

Low Intensity Magnetic separation (LIMS) was investigated ahead of the leach circuit feed as a pretreatment, to reduce overall leach reagent consumption and optimise recovery.

The LIMS upgrade of magnetics from oxide feed and flotation tailings involves application of a 3,000 Gauss field to recovery magnetics. The magnetic separation step generated a magnetics stream, comprising predominantly magnetite. To upgrade this byproduct to produce a saleable iron concentrate (at a grade of ~65% Fe), an additional regrind and flotation circuit was required, primarily to reject the pyrrhotite.

A significant amount of testwork (~60 variability metallurgical samples/ composites and 1,141 pulp samples) was completed to quantify magnetic iron within the deposit, which was completed successfully. The work identified a potential iron concentrate byproduct which can be produced with an average mass pull of ~5% across all sulphide and oxide composites.

14.1.2 Logistics and marketing

The iron concentrate (predominantly magnetite) produced in testwork has been assessed by two leading iron technical and marketing specialist consultants, who have formed a view that the product is commercially attractive to a range of specialist steel mills.

Assays of concentrates produced to date indicate low levels of deleterious elements with an expected iron concentrate specification shown in Table 33.

Table 33. Iron concentrate specifications.

Metal	Unit	Expected range
Iron	%	65
MgO	%	0.8-3
SiO2	%	2.1-3.4
Al2O3	%	0.2-0.8
TiO2	%	0.1-0.3
Manganese	%	0.1-0.4

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Metal	Unit	Expected range
CaO	%	0.03-0.10
Phosphorus	ppm	40-80

The iron concentrate is expected to be able to be sold to steel mills in Asia, where it would receive the most favourable offtake terms. Base offtake assumptions, as informed by marketing specialists, are assumed to be the Platts 62% Iron Ore index (CFR China), with a premium of US$3.40/t to reflect its 65% Fe content.

The product is expected to have a P80 of ~20µm which is best suited to pelletising or briquetting for end-user application. Pelletising is a well-established process used in the iron and steel industry. The process converts ultra fine concentrate to a product that can be fed direct into blast or electric furnaces. Briquetting is not commonly used in conventional blast furnace operations but is used in electric arc furnace applications.

14.1.3 Economic analysis

Logistics cost estimates, plant capital cost estimates and indicative marketing assumptions were determined in order to derive the value implications of the new potential byproduct. The analysis showed:

« An incremental ~A$180M capital investment is estimated in Stage 2 (at 14Mtpa process throughput rate) to establish the pyrrhotite flotation rejection circuit, concentrate handling facilities and new concentrate pipeline to a suitable facility (near Muchea).
« An incremental operating cost of ~A$1.10/t processed.
« Transport costs of A$80-110/wmt of concentrate to Asia.

The analysis completed showed that the opportunity is worthy of future assessment as part of the Stage 2 expansion, and is expected to be strongly incentivised when 62% Fe benchmark prices exceed ~US$110/t. It did not warrant inclusion in Stage 1, as the incremental value did not compensate for the additional capital investment required and additional complexity risk.

A potential iron concentrate byproduct could however provide an additional revenue stream further diversifying project revenues, which could be particularly valuable in periods of sustained commodity price lows. Producing a saleable iron concentrate could also increase mining inventory and reduce the strip ratio by converting marginal material to ore above economic cut-off, as well as increase the capacity of the TSF.

14.2 Resource/mining upside

The PFS has been constrained to a conventional open-pit mining phase, located entirely on Chaliceowned farmland. The current PFS open-pit mine plan only extracts ~50% of the Resource, leaving ~7.9Moz 3E, 450kt Ni, 250kt Cu, 46kt Co contained remaining in Resource under the modelled openpit. Upside beyond the modelled mining inventory therefore exists via:

« Extension of open-pit life and therefore a deeper open-pit, driven by higher mine design prices in ~2050.
« A potential transition to large scale underground mining (considered likely) – not currently modelled in the PFS.
« Growth in the Resource beyond the limit of drilling – the Resource extends to a depth of ~1.1km and high-grade mineralisation has been intersected up to ~900m beyond the limit of the Resource.

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==> picture [483 x 258] intentionally omitted <==

Figure 45. 3D view (looking east-north-east) of Gonneville block model (sulphide domains only).

The deepest mineralisation intersected to date is at ~1.1kmm below surface, where the host Gonneville Intrusion has a ~600m true thickness. The Intrusion and high-grade mineralisation remain open beyond this depth and hence Chalice considers Resource growth likely with further drilling.

Significant high-grade intersections from wide-spaced step-out drill holes beyond the current Resource include (refer to ASX Announcement on 23 April 2024):

« 5m @ 1.76 g/t 3E, 0.14% Ni, 0.09% Cu, 0.01% Co from 1164m (JD408);
« 10m @ 2.13g/t 3E, 0.18% Ni, 0.21% Cu, 0.01%% Co from 1191m (JD408 – the deepest mineralisation intersected at the Project to date);
« 10m @ 1.47g/t 3E, 0.14% Ni, 0.11% Cu, 0.01% Co from 845m(JD430);
« 7m @ 1.6g/t 3E, 0.19% Ni, 0.09% Cu, 0.02% Co from 1037m (JD430);
« 8m @ 1.14g/t 3E, 0.15% Ni, 0.13% Cu, 0.01% Co from 1052m (JD430);
« 6m @ 1.55g/t 3E, 0.15% Ni, 0.26% Cu, 0.01% Co from 1159m (JD430);
« 3.9m @ 2.94g/t 3E, 0.23% Ni, 0.55% Cu, 0.02% Co from 1197.1m (JD430);
« 17.1m @ 1.69g/t 3E, 0.16% Ni, 0.22% Cu, 0.02% Co from 1207m (JD430);
« 2m @ 7.08g/t 3E from 621m (JD431);
« 12m @ 1.48g/t 3E, 0.15% Ni, 0.14% Cu, 0.01% Co from 641m (JD431);
« 3.8m @ 1.6g/t 3E, 0.18% Ni, 0.14% Cu, 0.02% Co from 64.2m (JD432);
« 6m @ 1.81g/t 3E, 0.15% Ni, 0.08% Cu, 0.01% Co from 217m (JD433);
« 11.2m @ 1.43g/t 3E, 0.22% Ni, 0.12% Cu, 0.02% Co from 305.2m (JD433);
« 4m @ 1.11g/t 3E, 0.24% Ni, 0.11% Cu, 0.02% Co from 327m (JD433);
« 2m @ 1.44g/t 3E, 0.17% Ni, 0.02% Co from 620m (JD435);
« 3.2m @ 1.71g/t 3E, 0.23% Ni, 0.22% Cu, 0.02% Co from 765m (JD435);
« 6.4m @ 1.16g/t 3E, 0.14% Ni, 0.14% Cu, 0.01% Co from 412m (JD436);
« 12.6m @ 1.01g/t 3E, 0.21% Ni, 0.08% Cu, 0.02% Co from 442m (JD436);
« 2m @ 1.16g/t 3E, 0.2% Ni, 0.12% Cu, 0.02% Co from 457m (JD436); and,

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« 3m @ 2.26g/t 3E, 0.32% Ni, 0.15% Cu, 0.03% Co from 624m (JD436).

Given the scarcity of large, economic precious and base metal deposits, commodity prices (particularly precious metals prices) are expected to increase above the rate of inflation over the longer term. Hence, the mine design cut-off is expected to reduce over time, which would materially increase the mining inventory.

Increased commodity prices and mining inventory opens the possibility for further expansions of plant throughput capacity over the longer term (which would require additional approvals). In addition, the timing and scale of the Stage 2 plant expansion could be accelerated and enlarged if macroeconomic conditions incentivise.

Table 34. Mining opportunities and qualitative upside assessment.

Assessment of upside Mining opportunity potential Development of underground mining to recover Resources external to the currently modelled pit shells using bulk mining methods (including sub level or block caving) Higher long term prices converting material currently classified as waste to ore, and therefore enlarging the mining inventory and extending the life of the open-pit Open-pit optimisations (pit phasing, value-based cut-off, de-bottlenecking, blending, stockpiling and product mix optimisation, etc) Selective use of autonomous mining equipment (grade control drilling or haulage), to improve safety and operating efficiency Low grade mineralised waste on waste dumps becoming economic at higher commodity prices long term

14.3 Processing upside

The PFS has assessed several processing flowsheets, with the aim of maximising metallurgical recoveries whilst minimising costs and risk. Given the large scale of the Resource and unique characteristics of the Project, flowsheet design and optimisation will continue through the subsequent FS phase.

Further optimisations and recovery improvements are expected through a greater understanding of the geo-metallurgy as the FS progresses. Chalice will continue to assess opportunities for further downstream processing of the nickel concentrate with appropriate partners.

Processing trade-off studies and evaluations were completed at US$1,000/oz Pd. Spot prices are currently significantly higher than that of the trade-off studies, which incentive a flowsheet with higher input costs (e.g. grind size, reagent use, etc) to realise higher recoveries.

Table 35. Processing opportunities and qualitative upside assessment.

Assessment of Processing opportunity upside potential Commodity prices above mine design assumptions (US$1,000/oz Pd) incentivise a finer grind size and/or larger throughput, and higher reagent use to further increase metal recoveries and overall production Improved geo-metallurgical understanding and domaining of the deposit (spatially, mineralogically and metallurgically) potentially leading to improved recovery, Ni-CoPGM concentrate quality or ore-sorting removal of low-value material ahead of processing Enhanced metallurgical Ni-Co-Pt recoveries through flotation parameter optimisation, leaching optimisation, grind size optimisation (including staged grinding), and concentrate regrinding and grade optimisation

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Assessment of Processing opportunity upside potential Further downstream processing (offshore) of intermediate products with a vertically integrated partner to capture additional recovery and payability. New processing technologies (improved flotation cells, novel reagents, automation/machine learning in processing, comminution cost reductions etc) Technological improvements in comminution drive lower power usage and operating costs Additional revenue stream from monetising minor metals in concentrates (i.e. silver, rhodium, tellurium, etc) or sulphur (creating a valuable sulphuric acid byproduct in downstream processing)

14.4 Commercial upside

Preliminary discussions with offtake and finance partners indicated high levels of demand for concentrates sourced from safe western jurisdictions and a strong appetite to support projects like Gonneville.

Formal discussions with these partners will commence with the FS. As the scarcity of critical minerals becomes increasingly apparent across the western world, it is expected commercial terms will improve from the PFS assumptions.

Table 36. Commercial opportunities and qualitative upside assessment.

Assessment of Commercial opportunity upside potential Higher long-term realised metals prices / offtake terms due to scarcity of supply, lack of new large-scale discoveries particularly in stable jurisdictions and strong demand Strategic partnering to enhance offtake terms, and/or provide low-cost project finance or capital investment Co-operation, funding or incentives from third party sources (industry co-operatives, government grants / incentives, co-investment on infrastructure, etc) Offtake of product to non-smelter downstream processing facilities (reduced transportation costs) Additional payable metals in concentrate which are known to exist in the Deposit in minor quantities (Rh, Ir, Os, Ag, Te) Power price improvements through new PPAs and reductions in the SWIS Improvements in logistics costs through further engineering optimisations, synergies created via partnership opportunities and infrastructure upgrades (eg: road, rail, port)

14.5 Project risks

All mining projects have inherent risks that apply across the industry. Key risks specific to the Project are described in Table 37 with further work in the FS planned to mitigate risks where possible.

Table 37. Project Critical Risks

Category	Risk	Mitigation Approach
		Adherence to WA and Commonwealth regulations.
		Early and ongoing regulator engagement to understand
	Regulatory approval delays or	required study and environmental modelling deliverables
Permitting	design/operational constraints	that meet regulator expectations.
	enforced by regulators	Adherence to local and international standards.
		Clear demonstration of the application of the mitigation
		hierarchy to address potential environmental impacts.

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Category	Risk Mitigation Approach
	Significant investment in environmental offsets and biodiversity projects.
	Loss of social licence or disruption to construction/operations Stakeholder Engagement Plan developed, resourced and implemented, with ability to adapt quickly to increased public interest and changing regulatory requirements. Active consultation, presence and significant investment in local community since discovery in 2020. Community concerns addressed through transparent communications and responsiveness. Adherence to WA and Commonwealth regulations. Demonstration of the significant long term economic benefits of the Project through employment, servicing and royalties/taxation.
Operational	Resource model does not reconcile with grade control or production Deposit infill drilled to Indicated category (40m x 40m) to ~400m below surface. Four separate resource estimates performed to date on the deposit. Several different modelled approaches have been trialled, with negligible overall differences in tonnage or grade assessed. Localised 10m x 10m infill drilling completed to Measured resource category, with comparison completed vs the 40m x 40m estimate, with negligible tonnage or grade difference.
	Process plant does not perform as designed A very broad range of detailed testwork completed on 33 dedicated metallurgical holes, with significant rigour, technicl oversight and independent reviews. Over $15M invested to date in testwork and flowsheet design. Two separate and independent metallurgical labs used to complete testwork, with appropriate QAQC and cross checking of results. Piloting of RIP process in the FS to ensure continuous
	operation performance is as expected. This piloting is expected to underpin a Process Guarantee from an EPC provider. Flowsheet has three saleable products, providing a level of diversification.
	MDA access restriction or delays Logistics planning has been undertaken on several potential routes. Further stakeholder engagement will be undertaken to confirm the optimal routes to be used for each type of traffic.
	Strong collaboration with government agencies including Western Power, Water Corporation, Main Roads, Bunbury Port, Fremantle Port, Local shires of
	Supporting infrastructure Chittering and Toodyay to define suitable solutions.
	restrictions Potential for government funding pathways aligned to
	Major / Strategic Project status and the importance of the development of the project and provision of critical minerals to the State of WA and nationally.
	Hazardous materials and dust management result in unforeseen health & safety Adherence to WA Mining Guidelines Following best practice in industry, to develop operational procedures to minimise risks in mining and processing.

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Category	Risk Mitigation Approach
	restrictions and increased processing costs
	Deterioration in offtake terms Early and extensive engagement with potential offtakers. Initial indicative offtake terms received based on specifications of products from testwork. Concentrates produced from Gonneville are commercially standard and highly sought after,
	particularly from western sources. Lack of Ni-Cu-PGE discoveries worldwide means offtake terms expected to improve with increased concentrate deficits.
	Process water pipeline or grid connection delayed, or insufficient for Project requirements LOI in place with Water Corporation in relation to Alkimos water volumes required. Off-grid power solution scoped, which provides flexibility to commence operations without grid connection. Staging of the Project allows infrastructure to be constructed progressively. Extensive engagement with Water Corporation and Western Power.
Commercial /
	Allowance of contingency in capital estimates.
Economic
	Cost escalation Experienced procurement specialists to be used. Favourable Project location near Perth expected to drive highly competitive rates from service providers.
	Significant drop in commodity prices Mine design and PFS macro economic assumptions used are conservative, with positive cashflow margins expected even with a dramatically lower long term price environment Metals have historically had relatively uncorrelated prices providing some protection against unfavourable movements in multiple commodities at the same time. The scarcity of the metals found at Gonneville, particularly in the western world, means that commodity prices have historically increased in real terms over time. As geopolitical environments change, deposits of scale in stable jurisdictions become more difficult to find and extract, meaning commodity prices are expected to increase in real terms in the future.

15. Study team

Chalice utilised a specialist internal study team throughout the Gonneville PFS Study, with the support of several independent specialist consultants who Chalice would like to thank for their contribution. The following parties provided input into the study scopes:

« Study management
« Geology

Chalice

Chalice
« Resource modelling Cube Consulting (CP)
« Geotechnical
« Mining engineering
« Metallurgy
« Process engineering

Dempers & Seymour

Chalice, Entech Pty Ltd (CP)

Auralia Metallurgy, Strategic Metallurgy, A. Farghaly (CP) GR Engineering Services, NewPro, Ausenco

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« Tailings storage L&MGSPL, CMW, AQ2, Geoanalytica « Power supply Western Power, ECG Engineering, Accenture « Transport and logistics Qube Logistics, NMT Logistics « Waste Characterisation Graeme Campbell & Associates, SRK « Water supply Water Corporation, Fortin Pipelines « Environmental Biologic, Syrinx, GHD « Heritage and native title Dortch Cuthbert, Garwood Consulting « Social and community Voconiq « Risk, health and safety Chalice
« Risk, health and safety
« Marketing Rob Aird, John Clout, Warwick Davies
« 3[rd] party reviewers Enthalpy

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16. Appendices

A-1 Metallurgical Recovery Technical Data

A-1-1 Sulphide ore mass balances

The metallurgical recovery data presented in the tables below were used to generate the recovery algorithms used for the PFS, which apply a metal recovery to concentrate to each block within the diluted block model according to its grade.

The flotation testwork recovery data was adjusted to develop a regression targeting a 20% Cu and 8% Ni concentrate, using established relationships between recovery and concentrate grade.

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Table 38. HG2 Yr1-4 38μm

Product	Weight	Cu		Ni		S			Fe	Fe			Pd	Pd		Pt	Pt			Au				MgO	MgO		Co
	%	%	%rec	%	%rec	%	%rec		%		%rec		g/t		%rec	g/t		%rec		g/t		%rec		%	%rec		%		%rec
Head Assay		0.23		0.27		2.26			11.8				1.02			0.21				0.02				28.6			0.03
Cu Conc	0.66	27.6	84.3	1.40	3.50	32.5	9.47		30.9		1.70		58.3		35.6	2.56		8.40		1.49		42.0		2.39	0.06		0.11		2.93
Ni Conc	1.56	0.51	3.66	8.82	52.0	42.0	29.0		40.5		5.26		14.5		21.0	3.18		24.7		0.25		17.0		1.87	0.10		0.78		49.4
Leach					3.77										26.1			<10.00				6.63
Overall Recovery			84.3		55.8										82.8			33.1				65.6							49.4
Table 39. HG4	Yr1-4 38μm
Product	Weight	Cu		Ni		S			Fe				Pd			Pt			Au				MgO			Co
	%	%	%rec	%	%rec	%	%rec		%		%rec		g/t		%rec	g/t		%rec	g/t		%rec		%		%rec	%		%rec
Head Assay		0.21		0.24		1.73			10.8				0.83			0.14			0.03				29.1			0.02
Cu Conc	0.85	19.8	82.3	0.63	2.37	24.7	12.3		26.6		2.06		37.5		36.0	1.49		7.47	1.49		35.3		9.46		0.27	0.04		1.67
Ni Conc	1.47	0.82	5.93	8.43	54.9	38.7	33.4		36.6		4.91		16.0		26.6	4.79		41.6	0.57		23.5		3.75		0.19	0.75		50.0
Leach					2.47										20.2			5.48			31.4
Overall Recovery			82.3		57.4										82.7			54.5			90.2							50.0
Table 40. HG2	Yr5+ 38μm
Product	Weight	Cu		Ni		S				Fe				Pd			Pt			Au				MgO			Co
	%	%	%rec	%	%rec	%		%rec		%		%rec		g/t	%rec		g/t	%rec		g/t		%rec		%	%rec		%		%rec
Head Assay		0.23		0.20		1.48				11.2				1.09			0.26			0.07				31.0			0.02
Cu Conc	0.69	25.5	76.7	0.86	3.14	29.7		15.1		28.2		1.67		67.4 43.6			7.41 19.5			7.36		70.0		5.78	0.13		0.08		3.31
Ni Conc	0.96	0.97	4.09	7.94	40.5	27.1		19.2		26.9		2.23		12.6 11.4			5.73 21.1			1.13		15.0		10.6	0.33		0.78		48.0
Leach					<10.00										25.9			18.5				12.2
Overall Recovery			76.7		40.5										81.0			59.1				97.2							48.0

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Table 41. HG4 Yr5+ 38μm

Product	Weight	Cu		Ni		S		Fe		Pd		Pt		Au		MgO		Co
	%	%	%rec	%	%rec	%	%rec	%	%rec	g/t	%rec	g/t	%rec	g/t	%rec	%	%rec	%	%rec
Head Assay		0.13		0.17		0.95		10.1		0.83		0.16		0.04		32.6		0.02
Cu Conc	0.45	22.3	72.7	0.97	2.61	27.1	12.2	26.3	1.12	60.5	30.9	2.43	5.62	4.09	43.4	8.10	0.12	0.08	2.74
Ni Conc	1.04	1.22	9.21	6.99	43.5	34.6	36.3	32.0	3.16	21.4	25.5	7.51	40.5	1.17	28.8	7.22	0.24	0.70	54.7
Leach					3.62						22.5		10.8		19.5
Overall Recovery			72.7		47.1						78.9		56.9		91.7				54.7

Table 42. LG CR2 Nov 38μm

Product	Weight Cu Ni S Fe Pd Pt Au MgO Co
% % %rec % %rec % %rec % %rec g/t %rec g/t %rec g/t %rec % %rec % %rec
Head Assay	0.10 0.17 1.19 10.9 0.58 0.15 0.01 32.0 0.02
Cu Conc 0.52 17.0 83.0 0.84 2.72 21.2 9.15 21.2 1.02 39.7 31.0 1.90 5.82 1.03 44.4 13.0 0.21 0.09 2.67
Ni Conc	0.74 0.65 4.56 7.74 35.9 34.4 21.3 32.6 2.24 15.7 17.6 6.56 28.7 0.23 14.1 7.36 0.17 0.92 40.7
Leach 2.61 26.9 3.86 36.2
Overall Recovery 83.0 38.5 75.5 38.4 94.7 40.7

Table 43. LG S21 38μm

Product	Weight	Cu		Ni		S		Fe		Pd		Pt		Au		MgO		Co
	%	%	%rec	%	%rec	%	%rec	%	%rec	g/t	%rec	g/t	%rec	g/t	%rec	%	%rec	%	%rec
Head Assay		0.07		0.16		1.03		10.1		0.59		0.12		0.05		32.5		0.02
Cu Conc	0.24	25.3	70.4	0.81	1.21	30.2	7.53	27.6	0.62	51.8	22.2	3.71	7.48	1.75	22.9	5.15	0.04	0.09	1.16
Ni Conc	0.81	0.97	9.19	7.82	40.2	33.6	28.6	32.4	2.51	12.9	19.0	4.18	28.9	0.85	38.2	7.00	0.18	0.85	39.3
Leach					2.32						31.0		4.01		18.8
Overall Recovery			70.4		42.5						72.3		40.3		80.0				39.3

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Table 44. LG PYX C2 38μm (*open cycle nickel test)

Product	Weight	Cu		Ni		S		Fe		Pd		Pt		Au		MgO		Co
	%	%	%rec	%	%rec	%	%rec	%	%rec	g/t	%rec	g/t	%rec	g/t	%rec	%	%rec	%	%rec
Head Assay		0.15		0.15		0.86		10.9		0.65		0.19		0.05		31.5		0.02
Cu Conc	0.34	25.2	62.9	0.31	0.67	28.9	13.0	26.5	0.87	26.3	14.7	0.83	1.86	4.86	35.7	5.95	0.06	0.03	0.59
Ni Conc*	0.68	0.72	3.57	6.15	26.9	27.2	24.5	27.8	1.82	12.8	14.3	5.07	22.6	0.64	9.43	12.1	0.26	0.69	28.9
Leach					2.69						35.8		<10.00		41.7
Overall Recovery			62.9		29.6						64.8		24.4		86.8				28.9

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A-2 Competent Person Statement

A-2-1 Mining and Reserves

The information in this announcement that relates to Ore Reserves in relation to the Gonneville Project is based on and fairly represents information and supporting documentation compiled or reviewed by Dan Donald.

Mr Donald a full-time employee of Entech Pty Ltd and is a Member of the Australasian Institute of Mining and Metallurgy. Mr Donald has sufficient experience that is relevant to the style of mineralisation and type of deposit under consideration and the activity being undertaken to qualify as a Competent Person as defined in the JORC Code 2012 Edition. Mr Donald consents to the inclusion in the announcement of the matters based on his information in the form and context in which it appears.

A-2-2 Metallurgy

The information in this announcement that relates to metallurgical testwork results in relation to the Gonneville Project is based on and fairly represents information and supporting documentation compiled by Mr Adam Farghaly, BSc Eng, who is the Lead Metallurgist for the Company. Mr Farghaly is a Competent Person, and a Member of the Australasian Institute of Mining and Metallurgy. He is a qualified metallurgist and has sufficient experience that is relevant to the activity being undertaken to qualify as a Competent Person as defined in the 2012 edition of the Australasian Code for Reporting of Exploration Results, Minerals Resources and Ore Reserves. Mr Farghaly holds performance rights in Chalice Mining Limited. He consents to the inclusion in the announcement of the matters based on his information in the form and context in which it appears.

A-2-3 Exploration Results

The information in this announcement that relates to previously reported exploration results is extracted from the following ASX announcements:

« “New wide high-grade zones in ~900m step-out drill hole”, 31 July 2023.
« “High-grade copper-PGE zones extended at Gonneville”, 30 November 2023.
« “Gonneville Resource Remodelled to Support Selective Mining”, 23 April 2024.
« “Gold-copper Exploration Strategy for the West Yilgarn”, 3 September 2024.
« “Major metallurgical breakthrough at Gonneville”, 17 February 2025
« “Further process flowsheet improvements at Gonneville”, 6 May 2025

The above announcements are available to view on the Company’s website at www.chalicemining.com. The Company confirms that it is not aware of any new information or data that materially affects the exploration results included in the relevant original market announcement. The Company confirms that the form and context in which the Competent Person’s findings are presented have not been materially modified from the relevant original market announcement.

A-2-4 Mineral Resources

The Mineral Resource estimate referred to in this announcement was first announced by the Company in accordance with ASX Listing Rule 5.8 in its announcement dated 23 April 2024 titled “ Gonneville Resource remodelled to support selective mining ”. The Company confirms that it is not aware of any new information or data that materially affects the information included in the previous announcement and that all material assumptions and technical parameters underpinning the estimate in the previous announcement continue to apply and have not materially changed.

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A-2-5 Resource estimation methodology

All geological wireframe interpretations used in the Resource were constructed by Chalice using a combination of Leapfrog and Micromine software. Geological wireframes provided by Chalice include weathering, lithological, litho-geochemical and supergene/dispersion zone interpretations. Block modelling and grade estimation was carried out by Cube Consulting using Surpac, Datamine and Isatis software.

Prior to estimation, variables with below detection limit assays were assigned a positive value equal to half of the detection limit for the relevant grade variable. Intentionally unsampled intervals were retained as absent grade values. The vast majority of the intentionally unsampled intervals occur outside of the host intrusion lithology, and therefore have no bearing on the grade estimates.

Density is generally more poorly informed than the elemental variables, due to only core being sampled for density, but it was deemed possible to fill in unsampled density values on the basis of a multi-linear regression of sampled density values against the well-correlated and more widely informed Co, Fe, Ni and S variables.

All wireframes and drill data were rotated 40° anti-clockwise and placed in a local grid for estimation and mining studies. This brings the average strike of the mineralisation approximately in line with the local north-south axis.

All drillhole samples were flagged according to the geological and mineralisation domain interpretations provided by Chalice. Sample populations were statistically analysed to derive geostatistical domain groupings for Pd, Pt, Ni, Co, Cu, Au, As, S, Mg, Cr and density. Statistical analysis included comparison of global grade distributions, derivation of statistical correlations between grade variables and contact analysis of grade variables across the various geological domains. From analysis, estimation domains were determined for Pd/Pt, Ni/Co, Cu/Au, As, S, Mg, Cr and density variable groupings.

For primary high grade Pd, Pt, Ni and Co, mineralisation located within the Ultramafic intrusion, grade interpolation was undertaken using Ordinary Kriging (OK). For the high grade Cu/Au grouping, a mix of OK and Localised Uniform Conditioning (LUC) was used. For all six economic elements and S, the lower grade material outside of the high grade zones, situated within the general Ultramafic zone, was estimated using LUC. The lower grade general Ultramafic zone was divided into a low-tomoderate grade “Main” sub-domain, and very low-grade northwest sub-domain for Pd, Pt, Ni, Co, Cu, Au and S.

OK estimates for the granite, gabbro, and sediment lithologies were also undertaken, but using restrictive high-grade distance limiting parameters to curtail the propagation of rare high-grade samples. These high-grade samples are believed to be due mainly to re-mobilisation of mineralisation in the case of the surrounding sediments and granite. The mineralisation modelled outside of the Ultramafic envelope has not been classified as a Mineral Resource for reporting purposes.

For the secondary mineralisation, most notably in the supergene horizon, grade interpolation was undertaken using OK.

Indicator kriging was used to model the geometry of dyke material that was logged in the drill holes, typically represented by short and discontinuous intercepts, but which fell outside of the dyke Leapfrog wireframes. This additional dyke volume comprises approximately 1.4% of the total volume within the estimated Ultramafic intrusion envelope. Detection limit grades were assigned for all elemental variables and density was assigned based on density sample statistics within the dolerite dykes.

OK estimates were run into either 20mE x 20mN x 5mRL (local grid) parent blocks or 10mE x 20mN x 5mRL (local grid) parent blocks, which is approximately half the width of the nominal 40m infill drill spacing in the northing direction. Because of the north-south strike in local space, the nominally 60° easterly inclined drill holes, 1m downhole sample spacing and generally continuous nature of the

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variograms models for the economic elements, the local easting and RL block dimensions were set at a smaller 10m spacing. LUC estimates, where undertaken, were progressed to smaller 5mE x 10mN x 2.5mRL (local grid) blocks.

A variable variogram and search ellipse orientation strategy was implemented using Isatis’ DA functionality during grade interpolation to honour the local undulations in the mineralisation orientation. The hanging wall and footwall surfaces for the high grade zones were used to define the DA within the envelope of the Ultramafic intrusion. The Ultramafic contact was used for DA in the granite and sediment units. Constant rotations were used in the gabbro units, as these have relatively uniform dip and strike. The dyke hanging wall and footwall surfaces were used to inform the DA parameters for the estimation of the remaining dyke material not captured by wireframes. In the secondary zone, including the Supergene unit, the topographic, bottom of complete oxidation and top of fresh surfaces were used for DA.

Once estimation domains for grade interpolation were defined, composited drill hole sample populations were statistically analysed to derive grade capping values. It was observed that grade capping for the economic elements had an immaterial impact on the global grade. Boundary/contact analysis showed that the high grade mineralisation zones have hard boundaries with respect to the surrounding, lower-grade Ultramafic zone and so hard grade boundaries were applied to this contact. A general Ultramafic Main-NW sub-domain estimation boundary was also defined for Pd, Pt, Ni, Co, Cu, Au and sulphur interpolation, based on a large change in the grade distribution, and was treated as soft during interpolation, although different capping, variogram and search parameters were implemented either side of this boundary.

Final block values for Pd, Pt, Ni, Co, Cu, Au, S, Mg, Cr and density were validated by way of visual review of plans and cross sections (block model and drill samples presented with same colour legend), swath plots, and comparison of estimation domain mean grades with the input grade distribution data.

A-2-6 Forward Looking Statements

This announcement includes forward looking statements that have been based on an assessment of present economic and operating conditions, and assumptions regarding future events and actions that, as at the date of this announcement, are considered reasonable by the Company. Such forward-looking statements are not guarantees of future performance and involve known and unknown risks, uncertainties, assumptions and other important factors, many of which are beyond the control of the Company and its Directors and management. The Company cannot and does not give any assurance that the results, performance or achievements expressed or implied by the forward-looking statements will actually occur and investors are cautioned not to place undue reliance on these forward-looking statements. The Company has no intention to update or revise forward-looking statements, except where required by law.

A-2-7 Non-IFRS Financial Measures

The Company uses certain financial measures, such as net present value (NPV) and internal rate of return (IRR), to assess the projected performance of the Project. These measures, collectively referred to as Non-IFRS Financial Measures, are not recognised under International Financial Reporting Standards (IFRS). While the Company believes these metrics provide useful insight into the estimated financial outcomes derived from the PFS, they should not be considered in isolation or as substitutes for performance or cash flow measures prepared in accordance with IFRS.

As the financial forecasts and economic analysis contained in this announcement are not based on IFRS, the Non-IFRS Financial Measures presented do not have standardised definitions. Accordingly, the methods used to calculate these measures may differ from those applied by other companies, and therefore may not be comparable to similarly titled measures reported elsewhere. Investors should exercise caution when relying on these Non-IFRS Financial Measures.

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A-3 JORC Tables

A-3 JORC Tables	A-3 JORC Tables
A-3-1 Section 1: Sampling Techniques and Data
Criteria	JORC Code explanation Commentary
Sampling techniques	Nature and quality of sampling (e.g. cut channels, random chips, or specific specialised industry standard measurement tools appropriate to the minerals under investigation, such as down hole gamma sondes, or handheld XRF instruments, etc). These examples should not be taken as limiting the broad meaning of sampling. PQ diamond core samples were obtained for the development of the composites and samples used in the metallurgical test work. Mineralised zones were identified through analysis of, and comparison with, pre- existing assays from adjacent twin holes, XRF instrumentation and visual identification of mineralisation through geological logging.
	Include reference to measures taken to ensure Samples for metallurgical test work were selected from mineralised zones throughout the deposit that best represented the variable ore types. Sample intervals sourced for metallurgical test work from JDMET012 to JDMET028
	sample representivity and the appropriate (Phase 12) were selected through analysis
	calibration of any measurement tools or of, and comparison with, pre-existing assays
	systems used. from adjacent twin holes, XRF scan analysis
	and visual identification of mineralisation through geological logging. Sample intervals from JDMET029 to JDMET033 (Phase 13) were selected using assays from quarter core which were sent for analysis.
	Aspects of the determination of mineralisation that are Material to the Public Report. In cases where ‘industry standard’ work has been done this would be relatively simple (e.g. ‘reverse circulation drilling was used to obtain 1 m samples from which 3 kg was pulverised to produce a 30 g charge for fire assay’). In other cases more explanation may be required, such as where there is coarse gold that has inherent sampling problems. Unusual commodities or mineralisation types (e.g. submarine nodules) may warrant disclosure of detailed information. For the sample intervals sourced from JDMET012 to JDMET028 (Phase 12), mineralisation is recognised by the presence of sulphides within the host Ultramafic rock. In diamond core, sample intervals were selected on a qualitative assessment of the geology and sulphide content, compared with the results of XRF scan analysis and the results of pre-existing assays from adjacent twin holes. For sample intervals selected from JDMET029 to JDMET033 (Phase 13), mineralisation is recognised by the presence of sulphides within the host Ultramafic rock as well as from the quarter core drill assays.
	Drill type (e.g. core, reverse circulation, open- Diamond drill core is PQ size (85mm diameter). Triple tube has been used from
	hole hammer, rotary air blast, auger, Bangka, surface until competent bedrock and then
Drilling	sonic, etc) and details (e.g. core diameter, triple or standard tube, depth of diamond tails, face-sampling bit or other type, whether core standard tube thereafter. PQ is drilled at a maximum of 3m runs.
techniques

	is oriented and if so, by what method, etc). Core orientation is by an ACT Reflex (ACT III RD) tool
Drill sample recovery	Method of recording and assessing core and chip sample recoveries and results assessed. Individual recoveries of diamond drill core samples were assessed quantitively by comparing measured core length with expected core length from drillers mark. Generally, core recovery was excellent in fresh rock and approaching 100%. Core recoveryinoxidematerial is oftenpoordue

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Criteria	JORC Code explanation Commentary
	to sample washing out. Core recovery in the oxide zone averages 60%
	Diamond drilling triple tube coring in the
	Measures taken to maximise sample recovery oxide zone was undertaken to improve core
	and ensure representative nature of the sample recovery. This results in better core
	samples. recoveries but recovery is still only moderate
	to good.

	Whether a relationship exists between sample recovery and grade and whether sample bias may have occurred due to preferential loss/gain of fine/coarse material. There is no evidence of a sample recovery and grade relationship in unweathered material. Paired statistical analyses comparing AC, RC and DD samples from throughout the deposit show that there is no statistically significant difference between these sample types. RC grades are observed to be slightly higher than DD grades, but mostly in the <0.1ppm Pd range, resulting in an immaterial impact on the global resource. All three sample types were therefore considered compatible for use in the grade interpolation.
	Whether core and chip samples have been geologically and geotechnically logged to a level of detail to support appropriate Mineral Resource estimation, mining studies and metallurgical studies. All drill holes were logged geologically including, but not limited to; weathering, regolith, lithology, structure, texture, alteration and mineralisation. Logging was at an appropriate quantitative standard for metallurgical sample selection.
Logging
	Whether logging is qualitative or quantitative in nature. Core (or costean, channel, etc) photography. Logging is considered qualitative in nature. Diamond drill core is photographed wet before cutting.

	The total length and percentage of the relevant intersections logged. All holes were geologically logged in full.
Sub-sampling techniques and sample preparation	If core, whether cut or sawn and whether quarter, half or all core taken. Sample intervals selected for test work from JDMET012 to JDMET028 (Phase 12) comprised diamond core samples in their entirety to provide sufficient sample volume. Sample intervals selected for test work from JDMET029 to JDMET033 (Phase 13) comprised three quarters (¾) of the PQ diamond core. Samples, typically comprising 10-12m lengths of full core, were crushed in their entirety and then sub-sampled at the metallurgical laboratory. None of these samples are being used for Resource estimation or similar purposes.
	If non-core, whether riffled, tube sampled,
	rotary split, etc and whether sampled wet or Diamond core only.
	dry.
	For all sample types, the nature, quality and appropriateness of the sample preparation technique. Sample preparation is industry standard and comprises jaw crushing and sub-sampling for separate testing requirements at different crush sizes.

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Criteria	JORC Code explanation Commentary
	Quality control procedures adopted for all sub-sampling stages to maximise representivity Not applicable to metallurgical samples
	of samples.
	Measures taken to ensure that the sampling is representative of the in-situ material collected, including for instance results for field duplicate/second-half sampling. In all cases the entire length of core has been sampled and assayed as a single interval.
	Drill sample sizes are considered
	Whether sample sizes are appropriate to the appropriate for the style of mineralisation
	grain size of the material being sampled. sought and the nature of the drilling
	program.
Quality of assay data and laboratory tests	The nature, quality and appropriateness of the assaying and laboratory procedures used and whether the technique is considered partial or total. Pre-existing diamond drill core samples that were twinned as part of the metallurgical drill campaign underwent sample preparation and geochemical analysis by ALS Perth. Au-Pt-Pd was analysed by 50g fire assay fusion with an ICP-AES finish (ALS Method code PGM-ICP24). A 34 element suite was analysed by ME-ICP (ALS method code ME-ICP61) including Ag, Al, As, Ba, Be, Bi, Ca, Cd, Co, Cr, Cu, Fe, Ga, K, La, Mg, Mn, Mo, Na, Ni, P, Pb, S, Sb, Sc, Sr, Th, Ti, Tl, U, V, W, Zn, Zr. Additional ore-grade analysis was performed as required for elements reporting out of range for Ni, Cr, Cu (ALS method code ME-OG-62) and Pd, Pt (ALS method code PGM-ICP27). These techniques are considered total digests. Assays for the metallurgical testwork have been undertaken by Nagrom using similar methods as described above.
	For geophysical tools, spectrometers,
	handheld XRF instruments, etc, the parameters used in determining the analysis including instrument make and model, reading times, calibrations factors applied and their Not applicable as no such tools or instruments were used for the assay of metallurgical composites.
	derivation, etc.
	Nature of quality control procedures adopted (e.g. standards, blanks, duplicates, external laboratory checks) and whether acceptable levels of accuracy (i.e. lack of bias) and precision have been established. Certified analytical standards, blanks and duplicates were inserted at appropriate intervals for diamond, RC and AC drill samples with an insertion rate of >10%. Approximately 5% of >0.1g/t Pd assays were sent for cross laboratory checks. All QAQC samples display results within acceptable levels of accuracy and precision.
	Significant drill intersections are checked by
	The verification of significant intersections by either independent or alternative company personnel. the Project Geologist and then by the Exploration Manager. Significant intersections are cross-checked with the logged geology and drill core after final
Verification of
sampling and	assays are received.
assaying	The use of twinned holes. All samples obtained for metallurgical test work have been drilled as twin holes of pre- existing diamond holes within the Mineral Resource Estimate area and provide a comparisonbetweengrade/thickness

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Criteria	JORC Code explanation Commentary
	variations over a maximum of 5m separation between drill holes.
	Primary drill data was collected digitally using OCRIS software before being
	Documentation of primary data, data entry transferred to the master SQL database.
	procedures, data verification, data storage All procedures including data collection,
	(physical and electronic) protocols. verification, uploading to the database etc
	are captured in detailed procedures and summarised in a single document.
	Discuss any adjustment to assay data No adjustments were made to the lab reported assay data.
	Accuracy and quality of surveys used to locate drill holes (collar and down-hole surveys), trenches, mine workings and other locations used in Mineral Resource estimation. Diamond drill hole collar locations are recorded by Chalice employees using a handheld GPS with a +/- 3m margin of error.
Location of
data points	Specification of the grid system used. The grid system used for the location of all drill holes is GDA94 - MGA (Zone 50).

	Quality and adequacy of topographic RLs for reported holes were derived from
	control. handheld GPS pick-ups.
Data spacing and distribution	Data spacing for reporting of Exploration Results. Not applicable – only new metallurgical testwork results being reported.
	Whether the data spacing and distribution is sufficient to establish the degree of geological and grade continuity appropriate for the Not applicable. No drilling results reported and no new Mineral Resource Estimate is being reported.
	Mineral Resource and Ore Reserve estimation procedure(s) and classifications applied. Samples for metallurgical test work have been selected from holes throughout the deposit.
	Whether sample compositing has been applied. Metallurgical samples were composited from contiguous lengths of drill core as selected as described above.
	Diamond holes drilled to obtain sample for
	Whether the orientation of sampling achieves metallurgical test work were twins of pre- existing diamond holes that form part of the
	unbiased sampling of possible structures and the extent to which this is known, considering the deposit type. Resource. Original drill holes were typically oriented within 15° of orthogonal to the interpreted
Orientation of
data in relation
to geological	dip and strike of the zone of mineralisation.
structure	If the relationship between the drilling orientation and the orientation of key mineralised structures is considered to have introduced a sampling bias, this should be assessed and reported if material. The orientation of the drilling is not considered to have introduced sampling bias.

	Diamond core samples were collected in appropriately sized core trays and, following
Sample
	The measures taken to ensure sample security. orientation and mark-up, were submitted to Auralia Metallurgy by a Chalice contractor
security
	where they were processed and composited.
Audits or reviews	The results of any audits or reviews of sampling techniques and data. No audit of the metallurgical sampling has been completed. Metallurgical assays have been checked by a round robin analysis using 4 different laboratories. Metallurgical testwork isvalidated by back

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Criteria	JORC Code explanation	Commentary
		calculating and comparing the head
		assays from the tails and concentrate
		assays.

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A-3-2 Section 2: Reporting of Exploration Results

Criteria	JORC Code explanation Commentary
Mineral tenement and land tenure status	Type, reference name/number, location and ownership including agreements or material issues with third parties such as joint ventures, partnerships, overriding royalties, native title interests, historical sites, wilderness or national park and environmental settings. The Project comprises exploration licences E70/5118, E70/5119 and E70/5353, which are in good standing. The holder CGM (WA) Pty Ltd is a wholly owned subsidiary of Chalice Mining Limited. There are no known encumbrances.
	All drilling has occurred on granted
	The security of the tenure held at the time of reporting along with any known impediments to obtaining a licence to Exploration Licences. There are no known impediments to obtaining a licence to operate.
	operate in the area. E70/5119 partially overlaps ML1SA, a State Agreement covering Bauxite mineral rights only.
Exploration done by other parties	Acknowledgment and appraisal of exploration by other parties. There was very limited exploration at Gonneville pre-Chalice. Chalice has compiled historical records dating back to the early 1960’s which indicate only three genuine explorers in the area, all primarily targeting Fe-Ti-V mineralisation. Over 1971-1972, Garrick Agnew Pty Ltd undertook reconnaissance surface sampling over prominent aeromagnetic anomalies in a search for ‘Coates deposit style’ vanadium mineralisation. Surface sampling methodology is not described in detail, nor were analytical methods specified, with samples analysed for V2O5, Ni, Cu, Cr, Pb and Zn, results of which are referred to in this announcement. Three diamond holes were completed by Bestbet Pty Ltd targeting Fe-Ti-V situated approximately 3km NE of JRC001 (the Gonneville discovery hole). No elevated PGE-Ni-Cu-Co assays were reported. Bestbet Pty Ltd undertook 27 stream sediment samples within E70/5119. Elevated levels of palladium were noted in the coarse fraction (-5mm+2mm). Finer fraction samples did not replicate the coarse fraction results. A local AMAG survey was flown in 1996 by Alcoa using 200m line spacing which has been used by Chalice for targeting purposes.
	The target deposit type is an orthomagmatic PGE-Ni-Cu-Co sulphide deposit, within the Yilgarn Craton. The style
	Deposit type, geological setting and style of mineralisation. of sulphide mineralisation intersected consists of massive, matrix, stringer and disseminated sulphides typical of
Geology

	metamorphosed and structurally overprinted orthomagmatic Ni sulphide deposits.
	A summary of all information material to the understanding of the exploration results Provided in the body of the text.

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Criteria	JORC Code explanation Commentary
Drill hole Information	including a tabulation of the following information for all Material drill holes: Easting and northing of the drill hole collar Elevation or RL (Reduced Level – elevation above sea level in metres) of the drill hole collar Dip and azimuth of the hole Down hole length and interception depth hole length.
	If the exclusion of this information is justified on the basis that the information is not
	Material and this exclusion does not detract No material information has been
	from the understanding of the report, the excluded.
	Competent Person should clearly explain why this is the case.
Data aggregation methods	In reporting Exploration Results, weighting averaging techniques, maximum and/or minimum grade truncations (e.g. Cutting of high grades) and cut-off grades are usually Material and should be stated. Not applicable – only new metallurgical testwork results being reported.
	Where aggregate intercepts incorporate short lengths of high-grade results and longer lengths of low grade results, the Not applicable – only new metallurgical
	procedure used for such aggregation should testwork results being reported.
	be stated and some typical examples of such aggregations should be shown in detail.
	The assumptions used for any reporting of metal equivalent values should be clearly stated. Not applicable – no metal equivalent values reported.
	These relationships are particularly important
	in the reporting of Exploration Results. Not applicable – only new metallurgical
Relationship	If the geometry of the mineralisation with respect to the drill hole angle is known, its nature should be reported. testwork results being reported.
between
mineralisation
widths and	If it is not known and only the down hole lengths are reported, there should be a clear statement to this effect (e.g. ‘down hole length, true width not known’). Not applicable – only new metallurgical testwork results being reported.
intercept lengths

	Appropriate maps and sections (with scales) and tabulations of intercepts should be
Diagrams	included for any significant discovery being Not applicable – no new exploration
	reported These should include, but not be discovery results reported.
	limited to a plan view of drill hole collar locations and appropriate sectional views.
Balanced reporting	Where comprehensive reporting of all Exploration Results is not practicable, representative reporting of both low and high grades and/or widths should be practiced to avoid misleading reporting of Exploration Results. Not applicable – no exploration results excluded and all metallurgical tests detailed which cover the full feed grade spectrum expected for a bulk open-pit mine.
	Other exploration data, if meaningful and material, should be reported including (but Other than the metallurgical results
Other substantive
exploration data	not limited to): geological observations; contained in this announcement, no new
	geophysical survey results; geochemical exploration results are reported.

	surveyresults;bulk samples – size and

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Criteria	JORC Code explanation Commentary
	method of treatment; metallurgical test results; bulk density, groundwater, geotechnical and rock characteristics; potential deleterious or contaminating substances.
Further work	The nature and scale of planned further work (e.g. Tests for lateral extensions or depth extensions or large-scale step-out drilling). No further resource definition drilling is envisaged at the Project prior to a potential Final Investment Decision. Some additional metallurgical sample and geotechnical drilling is being scoped as part of the Feasibility Study in CY26.
	Diagrams clearly highlighting the areas of possible extensions, including the main
	geological interpretations and future drilling Included within the report.
	areas, provided this information is not commercially sensitive.

A-3-3 Section 3: Estimation and Reporting of Mineral Resources

Criteria	JORC Code explanation Commentary
Database integrity	Measures taken to ensure that data has not been corrupted by, for example, transcription or keying errors, between its initial collection and its use for Mineral Resource estimation purposes. OCRIS data logging software is used by Chalice for front end data collection and has in-built validation for all geological logging and sampling. All logging, sampling and assay files are stored in a SQL Server database using DataShed (industry standard drill hole database management software). User access to the database is regulated by specific user permissions. Only the Database Manager can overwrite data. All data has passed a validation process; any discrepancies have been checked by Chalice personnel before being updated in the database.
	Cube Consulting completed validation checks on the drill hole data extraction provided by Chalice for use in the Mineral Resource Estimate. Multiple collar entries, potentially suspect collar and downhole survey results, absent survey or assay data, overlapping intervals, negative
	Data validation procedures used. sample lengths, out of range assay values and
	sample intervals which extended beyond the hole depth defined in the collar table were reviewed. Only minor validation issues were detected which were communicated to Chalice and corrected prior to the preparation of the Mineral Resource estimate.
Site visits	Comment on any site visits undertaken by the Competent Person and the outcome of those visits. A site visit to the Gonneville Project was completed by Mike Job (Principal Geologist/Geostatistician at Cube Consulting) and Mike Millad (Principal Geologist/Geostatistician at Cube Consulting) on 12 May2022, and an inspectionof theALS

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Criteria	JORC Code explanation Commentary
	sample preparation and analytical laboratories was undertaken by Mike Job on 2 June 2022. Mike Job and Mike Millad assume Competent Persons status for the Mineral Resource estimate. During the Gonneville site visit, the drilling, sampling, geological logging, density measurement and sample storage facilities, equipment and procedures were witnessed, and discussions held with Chalice representatives. The facilities and equipment were appropriate, and the procedures were well-designed and being implemented consistently. The sample preparation and analytical laboratories were well equipped and were operated to a very high standard. In the Competent Persons’ opinion, the geological and analytical data being produced is appropriate for use in a Mineral Resource estimate.
	If no site visits have been undertaken indicate why this is the case. Not applicable (see above).
Geological interpretation	Confidence in (or conversely, the uncertainty of) the geological interpretation of the mineral deposit. The location and orientation of the primary PGE- Ni-Cu-Co mineralisation within the Ultramafic host unit are reasonably well understood and have been developed over the course of the drill-out phase of the project. Geological controls on the supergene/dispersion zone material are reasonably simple and well understood. Confidence in the orientations of the barren Dolerite dyke lithology is variable over the footprint of the deposit, due to the geological complexity shown by this lithology unit. However, volumetrically the unit is considered as having been appropriately captured in the geological interpretation and by geostatistical interpolation of minor dolerite intervals not captured in the Leapfrog wireframes generated by Chalice. Work on improving definition of, and confidence in, the Dolerite lithology by Chalice is ongoing.
	Sample intercept logging and assay results from drill core and RC samples form the basis for the geological interpretations. A criterion of > 0.9ppm Pd has been used by Chalice to construct the supergene/dispersion zone mineralised wireframe. The logged oxide- transition boundary in the weathering profile
	Nature of the data used and of any was taken into account when developing the
	assumptions made. interpretation. A minimum intersection width of 2m was applied. High grade mineralisation wireframes were constructed separately for Pd, Cu and Ni using separate cut-off grades for each. The cut-off grades used were based on inflexions representing natural population breaks in the log probability plots. To preserve a level of continuitywhen interpretinghighergrade

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Criteria	JORC Code explanation Commentary
	zones, modelling allows a maximum of 1 hole with mineralisation below the cut-off grade between mineralised holes. i.e.: -If one hole with mineralisation below the cut-off grade is present between higher grade holes then the wireframe is pushed through the interpreted position using the minimum mining width. If two or more holes with mineralisation below the cut-off grade are present, the wireframe is not continued through the drillholes. Any high-grade intercepts which do not fit these criteria are not included in the wireframes and will instead be dealt with as part of the surrounding mineralised general Ultramafic zone. The high grade Pd zones were modelled first as the previous MRE#3 “G Zones” could be used to provide the general geometry. A mineralised intercept above a 0.9 Pd ppm cut off was calculated with the economic composite tool in LeapfrogGeo using the sulphide assay table. The intercept calculation allowed for a minimum ore composite length of 4m with a maximum 4m of internal waste and a maximum of 2m consecutive waste. The Pd intercepts were then classified using the interval select tool and finally domained using the vein tool. Sections were drawn viewing towards 40⁰N (NW-SE strike) for correlating the Pd zones. A mineralised intercept above a 0.18% Cu cut off was calculated with the economic composite tool in LeafrogGeo using the sulphide assay table. The intercept calculation allowed for a minimum ore composite length of 4m with a maximum 4m of internal waste and a maximum of 2m consecutive waste. The high grade Cu mineralisation could not be modelled in the same way as Pd as the intercepts were thicker but not as continuous from south to north through the Gonneville Ultramafic. Instead, the intrusion tool was utilised and the geometry based on the “nose” and the “embayment” models of MRE#3. A mineralised intercept above a 0.2% Ni cut off was calculated with the economic composite tool in LeafrogGeo using the sulphide assay table. The intercept calculation allowed for a minimum ore composite length of 4m with a maximum 4m of internal waste and a maximum of 2m consecutive waste. A mixture of the vein tool and intrusion tools were used to model the high grade Ni zones due to the varying geometry of mineralisation.
	The effect, if any, of alternative interpretations on Mineral Resource estimation. Alternative interpretations are likely to materially impact on the Mineral Resource estimate on a local, but not global, basis.
	The use of geology in guiding and The litho-geochemical domains within the host Ultramafic unit are known to have an association with the orientation of the primary
	controlling Mineral Resource estimation. mineralisation zones. The grades of the
	economic elements and geological interpretationsfor thesefeatureshave been

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Criteria	JORC Code explanation Commentary
	incorporated into the resource estimation approach via the development of trend surfaces informing a variable variogram and search ellipse orientation strategy (Dynamic Anisotropy (DA)).
	The factors affecting continuity both of grade and geology. The deposit represents part of a large layered intrusion. Sulphide content and metal grade are well correlated, with higher sulphide concentration generally corresponding to higher metal content within the Ultramafic intrusion. On a global scale the mineralisation displays good geological and grade continuity, which is largely governed by magmatic fractionation processes within the host intrusion. On a local scale geological and grade continuity is disrupted by the presence of variably oriented barren dolerite dykes and granite inclusions, both of which post-date and therefore overprint the mineralisation.
	The main part of the mineral resource within the Ultramafic extends for a strike length of approximately 1.8km and is 600 to 800 m thick. Plan width of the sub-parallel, high grade Pd
	The extent and variability of the Mineral Resource expressed as length (along zones ranges from m to ~60m, for the high grade Cu zones from 4m to ~160m and for the
Dimensions	strike or otherwise), plan width, and depth high grade Ni zones from m to ~60m. Plan width of the encompassing sulphide poor zones varies
	below surface to the upper and lower limits of the Mineral Resource. from 100 to 150m. The reported Measured Mineral Resource is within approximately 130m
	of surface. The reported Indicated Mineral Resource is within approximately 450m below surface. The reported Inferred Mineral Resource is within approximately 900m below surface.
Estimation and modelling techniques	The nature and appropriateness of the estimation technique(s) applied and key assumptions, including treatment of extreme grade values, domaining, interpolation parameters and maximum distance of extrapolation from data points. If a computer assisted estimation method was chosen include a description of computer software and parameters used. All geological and mineralisation wireframe interpretations used in the Resource were constructed by Chalice using a combination of Leapfrog and Micromine software. Geological wireframes provided by Chalice include weathering, lithological, litho-geochemical and supergene/dispersion zone interpretations. Block modelling and grade estimation was carried out by Cube Consulting using Surpac, Datamine and Isatis software. Statistical analysis was carried out by Cube Consulting using Geoaccess Professional and Isatis software. All wireframes and drill data were rotated 40° anti-clockwise and placed in a local grid for estimation and mining studies. This brings the average strike of the mineralisation approximately in line with the local grid north- south axis. Prior to estimation of variables, below detection limit assays were assigned a positive value equal to half of the detection limit for the relevant grade variable. Intentionally unsampled intervals were retained as absent grade values. The vast majority of the intentionally unsampledintervals occuroutside

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Criteria	JORC Code explanation	Commentary
		of the host intrusion lithology, and therefore
		have no bearing on the grade estimates.
		All drillhole samples were flagged according to
		the geological and mineralisation domain
		interpretations provided by Chalice. Sample
		populations were statistically analysed to derive
		geostatistical domain groupings for Pd, Pt, Ni,
		Co, Cu, Au, As, S, Mg, Cr and density. Statistical
		analysis included comparison of global grade
		distributions, derivation of statistical correlations
		between grade variables and contact analysis
		of grade variables across relevant geological
		domains. From analysis, estimation domains
		were determined for Pd/Pt, Ni/Co, Cu/Au, As, S,
		Mg, Cr and density variable groupings.
		Information regarding the in-situ mineral
		chemistry of the various mineral species for the
		deposit is currently not available. Mineral
		speciation was therefore not incorporated into
		the definition of the geostatistical domains.
		For primary high grade Pd, Pt, Ni and Co,
		mineralisation located within the Ultramafic
		intrusion, grade interpolation was undertaken
		using Ordinary Kriging (OK) For the high grade
		Cu/Au grouping, a mix of OK and Localised
		Uniform Conditioning (LUC) was used. For all six
		economic elements, the lower grade material
		outside of the high grade zones, situated within
		the general Ultramafic zone, was estimated
		using LUC. The lower grade general Ultramafic
		zone was divided into a low-to-moderate grade
		“Main” sub-domain, and very low-grade
		northwest sub-domain for Pd, Pt, Ni, Co, Cu and
		Au. OK estimates for the granite, gabbro, and
		sediment lithologies were also undertaken, but
		using restrictive high-grade distance limiting
		parameters to curtail the propagation of rare
		high-grade samples. These high-grade samples
		are believed to be due mainly to re-mobilisation
		of mineralisation in the case of the surrounding
		sediments and granite. The mineralisation
		modelled outside of the Ultramafic envelope
		has not been classified as a Mineral Resource
		for reporting purposes.
		Indicator kriging was used to model the
		geometry of dyke material that was logged in
		the drill holes, typically represented by short and
		discontinuous intercepts, but which fell outside
		of the dyke Leapfrog wireframes. This additional
		dyke volume comprises approximately 1.4% of
		the total volume within the estimated Ultramafic
		intrusion envelope. Detection limit grades were
		assigned for all elemental variables and density
		was assigned based on density sample statistics.
		Arsenic only occurs in very low abundances
		and was modelled using OK throughout. As is of
		higher grade in the southeast of the Ultramafic intrusion, and of lower grade to the north of this,
		hence a Main-SE subdivision was implemented.
		Sulphur was modelled using OK in the high
		grade domains andwith LUCin the surrounding

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		general Ultramafic. S estimation domains
		differed slightly from the economic elements, in
		that the litho-geochemical units were split
		about the top-of-fresh surface whereas the
		economic elements were split about the base
		of complete oxidation surface. The Main vs
		northwest domain subdivisions of the fresh
		Ultramafic zone was used for S modelling, similar
		to the economic elements. S was also
		interpolated using OK in the granite, gabbro,
		dyke and sediment lithologies, with appropriate
		high grade distance limits applied. It is
		noteworthy that in the immediate hanging wall
		and footwall of the Ultramafic intrusion, within
		the sediment lithological unit, S grades are
		elevated, which may have environmental
		implications for waste disposal.
		Mg and Cr were modelled using OK. It was
		observed that Mg is relatively depleted in the
		oxide zone while Cr is relatively enriched in the
		oxide zone and that there is no significant
		difference between Mg and Cr grades in the
		high grade mineralisation zones and
		surrounding general Ultramafic zones. A
		relatively simple domaining scheme was
		therefore used, whereby the general Ultramafic
		and high grade zones were rolled together into
		a single domain for estimation, with a split
		about the base of oxide surface.
		Density was modelled using OK within the
		transitional + fresh portion of the Ultramafic
		intrusion, granite, gabbro and sediment
		lithologies. Constant density assignments were
		made in the oxide zone, where the paucity of
		data did not justify using geostatistical
		interpolation. Density is generally more poorly
		informed than the elemental variables, due to
		only core being sampled for density, but it was
		deemed possible to fill in unsampled density
		values in the based on a multi-linear regression
		of sampled density values against the well-
		correlated and more widely informed Co, Fe, Ni
		and S variables, with which density is generally
		well correlated.
		All of the estimated variables were modelled
		independently using OK in the Supergene
		enrichment zone.
		Variogram models for Pd, Pt, Ni, Cu, Au, As, Cr,
		Mg and S were produced by first transforming
		the composite grades to Gaussian space in
		order to elucidate the true underlying spatial
		structure, before back-transforming to real
		space for use in interpolation. Ni and Co are
		strongly correlated and therefore the Ni
		variograms were used to interpolate Co. For the
		density variable, statistical and spatial variability
		is low within individual estimation domains, and hence variogram models could be produced
		directly in real space. The variography is
		generally characterised by strong anisotropy
		between the semi-major/majoraxis plane of

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		mineralisation (parallel to the tabular
		mineralised zones) and the perpendicular,
		shorter-range minor axis. Practical ranges for the
		main economic elements in the plane of
		mineralisation is generally of the order of 100m,
		while in the high-grade mineralisation zones it is
		most often between 40m and 50m. Variogram
		modelling was undertaken on capped grade
		values.
		Once estimation domains for grade
		interpolation were defined, composited drill
		hole sample populations were statistically
		analysed to derive grade capping values. It
		was observed that grade capping for the
		economic elements had an immaterial impact
		on the global grade. Boundary/contact analysis
		showed that the high grade mineralisation
		zones have hard boundaries with respect to the
		surrounding, lower-grade Ultramafic zone and
		so hard grade boundaries were applied to this
		contact. A general Ultramafic Main-NW sub-
		domain estimation boundary was also defined
		for Pd, Pt, Ni, Co, Cu, Au and S interpolation,
		based on a large change in the grade
		distribution, and was treated as soft during
		interpolation, although different capping,
		variogram and search parameters were
		implemented either side of this boundary. In
		addition to the grade caps, distance-based
		grade thresholds were also chosen and
		implemented for interpolation those zones
		where mineralisation is moderately or highly
		discontinuous (i.e. lower grade Ultramafic zones
		outside of the high grade domains, granite,
		gabbro, and sediment). This was based on
		observed inflexions in the grade histograms that
		are interpreted as representing the onset of the
		anomalous high grade sub-population. It is
		noted that the largely barren zones outside of
		the Ultramafic intrusion have not been classified
		as resources, and were modelled only to
		provide some indication in the block model of
		where these patches of mineralisation occur,
		and to show where sometimes high
		abundances of deleterious elements occur
		(e.g. high sulphur in the sediment footwall).
		Density bottom and top truncations have been
		applied, based on examination of density
		histograms, therefore completely excluding the
		outliers from the estimation process.
		Estimation of Pd, Pt, Ni, Co, Cu, Au, As, S, Mg
		and Cr was subsequently undertaken by OK for
		the primary and secondary mineralisation. As
		previously mentioned, the OK estimates were
		progressed to LUC estimates for Pd, Pt, Ni, Co,
		Cu, Au and S in the transitional + fresh portion of
		the Ultramafic intrusion outside of the high grade zones and in some of the larger Cu/Au high grade zones. Geostatistical interpolation of
		density was restricted to the transitional + fresh
		zones, with assignments being made in the
		oxidezone.A variablevariogramand search

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		ellipse orientation strategy was implemented
		using Isatis’ DA functionality during grade
		interpolation to honour the local undulations in
		the mineralisation orientation. The hanging wall
		and footwall surfaces for the high grade zones
		were used to define the DA within the envelope
		of the Ultramafic intrusion. The Ultramafic
		contact was used for DA in the granite and
		sediment units. Constant rotations were used in
		the gabbro units, as these have relatively
		uniform dip and strike. The dyke hanging wall
		and footwall surfaces were used to inform the
		DA parameters for the estimation of the
		remaining dyke material not captured by
		wireframes. In the secondary zone, including
		the Supergene unit, the topographic, bottom of
		complete oxidation and top of fresh surfaces
		were used for DA.
		Search and block plans were as follows:
		Primary mineralisation Pd, Pt, Ni, Co, Cu, and Au
		(within Ultramafic unit and high grade zones) –
		A minimum of 4 to 6 and maximum of 16
		samples per estimate into a parent block size of
		10 mI x 20 m(N) x 5 m(RL). The maximum limit
		was allowed to be exceeded in cases where
		samples are situated within any given block,
		since the condition was set whereby the OK
		would by default use all samples within the
		block. The maximum number of samples per
		drillhole was limited by using anisotropic
		distances for sample selection in combination
		with a maximum of 4 samples per search ellipse
		quadrant. A single search pass was used. Block
		discretisation scheme was 5 I(E) x 5 pts(N) x 2
		pts(RL). LUC post-processing of the six economic
		elements was into a Selective Mining Unit (SMU)
		block size I5 m(E) x 10 m(N) x 2.5 m(RL).
		Secondary mineralisation Pd, Pt, Ni, Co, Cu and
		Au (within the Ultramafic, high grade zones and
		Supergene unit)used a minimum of 4 to 6 and
		maximum of 16 samples per estimate into a
		parent block sizIf 10 m(E) x 20 m(N) x 5 m(RL).
		The maximum limit was allowed to be
		exceeded in cases where samples are situated
		within any given block, since the condition was
		set whereby the OK would by default use all
		samples within the block. The maximum number
		of samples per drillhole was limited by using
		anisotropic distances for sample selection in
		combination with a maximum of 4 samples per
		search ellipse quadrant. A single search pass
		was used. The block discretisation scheIwas
		pts(E) x 5 pts(N) x 2 pts(RL).
		For primary and secondary zones, S - A
		minimum of 4 to 6 and maximum of 16 samples
		per estimate into a parent bloIsize of 20 m(E) x
		20 m(N) x 5 m(RL). The maximum limit was allowed to be exceeded in cases where samples are situated within any given block,
		since the condition was set whereby the OK
		would by defaultuse allsampleswithin the

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Criteria	JORC Code explanation	Commentary
		block. The maximum number of samples per
		drillhole was limited by using anisotropic
		distances for sample selection in combination
		with a maximum of 4 samples per search ellipse
		quadrant. A single search pass was used. Block
		discretisatioIcheme was 5 pts(E) x 5 pts(N) x 2
		pts(RL). LUC post-processing of the S variable,
		where applicable, was into a Selective Mining
		Unit (SI block size of 5 m(E) x 10 m(N) x 2.5 m(RL).
		For the primary and secondary zone As, Cr and
		Mg, a minimum of 3 to 6 and maximum of 16
		samples per estimate were used into a pInt
		block size of 20 m(E) x 20 m(N) x
		5 m(RL). The maximum number of samples per
		drillhole was limited by using anisotropic
		distances for sample selection in combination
		with a maximum of 4 samples per search ellipse
		quadrant. A single search pass was used. High
		grade distance limiting was implemented in
		addition to grade capping in the largely barren
		units. The block discIisation scheme was 5 pts(E)
		x 5 pts(N) x 2 pts(RL).
		For the primary zone density within the
		Ultramafic intrusion, a minimum of 4 and
		maximum of 16 samples per estimate were used
		iI a parent block size of 5 m(E) x 10 m(N) x 2.5
		m(RL). Outside of the Ultramafic intruIn, a parent
		block size of 20 m(E) x
		20 m(N) x 5 m(RL) was used. The maximum
		number of samples per drillhole was limited by
		using anisotropic distances for sample selection
		in combination with a maximum of 4 samples
		per search ellipse quadrant. The maximum limit
		was allowed to be exceeded in cases where
		samples are situated within any given block,
		since the condition was set whereby the OK
		would by default use all samples within the
		block. A single search pass was used. The bIk
		discretisation scheme was 5 pts(E) x 5 pts(N) x 2
		pts(RL).
		For Pd, Pt, Ni, Co, Cu, Au, S, Mg and Cr, un-
		estimated domains (due to a paucity of
		samples) have been assigned constant grades
		based either on sample statistics or interpolated
		domain analogues. None of the ex-Ultramafic
		blocks, whether interpolated or assigned, have
		been classified as Mineral Resource.
		For As un-estimated blocks have been assigned
		half detection limit.
		For density, un-estimated blocks, inclusive of all
		secondary estimation domains, were assigned
		values based on applicable sample statistics.
		Final block values for Pd, Pt, Ni, Co, Cu, Au, S,
		Mg, Cr and density were validated by way of
		visual review of plans and cross sections (block
		model and drill samples presented with same colour legend), swath plots, and comparison of estimation domain mean grades with the input
		grade distribution data.

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Criteria	JORC Code explanation Commentary
	The Mineral Resource estimate was compared to the previous estimate undertaken by Cube Consulting in March 2023.
	The availability of check estimates, No previous mining has taken place at the
	previous estimates and/or mine project, and production data are not available
	production records and whether the to reconcile against the block model estimates.
	Mineral Resource estimate takes The Mineral Resource model has been peer
	appropriate account of such data. reviewed internally at Cube Consulting. Mr Mark
	Noppé of SRK undertook periodic high-level reviews of the estimation process on an in- stream basis of previous resource estimates.
	The assumptions made regarding recovery of by-products. Gonneville is a polymetallic deposit, and the assumption based on metallurgical testwork to date has been made that all reported constituents are recovered and are able to be sold.
	Sulphur, magnesium, chromium and arsenic have been estimated. As is observed to generally be of very low grade, while S is notably enriched in the immediate hanging
	Estimation of deleterious elements or other non-grade variables of economic significance (eg. sulphur for acid mine drainage characterisation). wall and footwall sediments of the Ultramafic intrusion, and especially so on the footwall side. Magnesium is observed to be relatively depleted in the oxide zone, while the opposite is true for chromium. No other deleterious variables have been estimated but to date there are no indications of any deleterious elements in concentrate samples.
	In the case of block model interpolation, the block size in relation to the average sample spacing and the search employed. OK estimates were run into either 10mE x 20mN x 5mRL or 20mE x 20mN x 5mRL (local grid) parent blocks, which is approximately half the width of the nominal 40m infill drill spacing in the northing direction. Because of the north-south orebody strike in local space, the nominally 60° easterly inclined drill holes, 1m downhole sample spacing and generally continuous nature of the variograms models for the economic elements, the local easting and RL block dimensions were set at a smaller 10m and 5m, respectively. LUC estimates, where undertaken, were progressed to smaller 5mE x 10mN x 2.5mRL (local grid) blocks.
	Within the Ultramafic unit the LUC modelling
	Any assumptions behind modelling of process for Pd, Ni, Cu, Au, Co, Pt and S has
	selective mining units. assumed an SMU size of 5 m E x 10 m N x 2.5 m
	RL.
	Any assumptions about correlation between variables. The high degree of observed correlation between Ni and Co grade meant that Ni variograms were used for Co interpolation. These elements are mostly bound together in pentlandite, hence the close relationship. Density was also observed to be well correlated with Ni, Fe, Co and S.
	Description of how the geological The litho- geochemical domains within the host
	interpretation was used to control the Ultramafic unit are known to have an
	resource estimates. association with the orientation of the primary mineralisation zones (i.e.thehighgrade

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Criteria	JORC Code explanation Commentary
	mineralisation zones). Geological interpretations for these features have been incorporated into the resource estimation approach via the development of trend surfaces informing a variable search ellipse orientation strategy (Dynamic Anisotropy). The geological interpretation for the supergene/dispersion zone has been used to constrain the resource estimate for The reported weathering zone material. a variable search ellipse orientation strategy (Dynamic Anisotropy) was employed to capture local undulations in the supergene/dispersion zone during grade estimation.
	Discussion of basis for using or not using grade cutting or capping. The need for grade capping was assessed for all estimated variables on a per estimation domain basis prior to estimation. Histograms and log-probability plots were used to review composited sample grade distributions graphically. Additionally, a visual inspection was carried out in Surpac for potential clustering of very high-grade sample data prior to selecting a capping value. Capping values, where deemed necessary, were applied to the composited sample grades. In addition to the grade caps, high grade distance limiting was implemented for high grade sub-populations in the largely barren domains and in the lower grade portion of the Ultramafic intrusion. Bottom and top truncations were applied to density composites on a per estimation domain basis.
	Final block values for Pd, Pt, Ni, Co, Cu, Au, As, S, Mg, Cr and density were validated by way of visual review of plans and cross sections (block model and drill samples presented with same
	The process of validation, the checking colour legend), swath plots, and comparison of
	process used, the comparison of model estimation domain mean grades with the input
	data to drill hole data, and use of grade distribution data. The block model
	reconciliation data if available. reflected the variability of the grades in the drillhole samples both globally and locally. No previous mining has taken place at the Project, and production data is not available to reconcile against the block model estimates.
Moisture	Whether the tonnages are estimated on a dry basis or with natural moisture, and the method of determination of the moisture content. Tonnages are estimated on a dry basis. No moisture data are available.
	Any oxide block within the optimisation pit shell above a Net Smelter Return (NSR) cut-off of A$25/t is considered as Mineral Resource
Cut-off	The basis of the adopted cut-off grade(s) amenable to mining by open pit methods.
parameters	or quality parameters applied. Any transitional or fresh block within the
	optimised pit shell above a NSR cut-off of A$25/t is considered as Mineral Resource amenable to mining by open pit methods.

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Criteria	JORC Code explanation	Commentary
		Any transitional or fresh block outside of the
		optimised pit shell, within a MSO shape and
		above a NSR cut-off of A$110/t is considered as
		Mineral Resource amenable to mining by
		underground methods.
		The determination of the NSR uses metal
		recovery assumptions and also incorporates
		assumptions relating to metal prices, metal
		payabilities, exchange rates, royalties, transport
		and treatment charges.
		For further information on the assumptions used
		in the NSR estimation refer to the Cut-off
		methodology section contained within the ASX
		Announcement on 23 April 2024.
		The Mineral Resource estimate is based on
	Assumptions made regarding possible mining methods, minimum mining dimensions and internal (or, if applicable,	conventional open cut drill, blast, load, and haul mining methods for the open pit portion of the resource.
	external) mining dilution. It is always	The pit optimisations prepared to support
	necessary as part of the process of	reasonable prospects for eventual economic
	determining reasonable prospects for	extraction had appropriate mining dilution and
Mining factors or assumptions	eventual economic extraction to consider potential mining methods, but the assumptions made regarding mining	ore loss applied. The Mineral Resource estimate itself is reported without mining dilution or ore loss.
	methods and parameters when estimating Mineral Resources may not always be rigorous. Where this is the case, this should be reported with an explanation of the basis of the mining assumptions made.	Consideration was given to the possibility of applying long hole open stoping underground mining methods to the sulphide resource outside of the optimised pit shell. Appropriate mining cost and commodity prices have been used to determine a cut-off grade for such an
		underground mining approach.
		Metallurgical test work for Resource reporting
		on oxide material conducted includes:
		Detailed QEMSCAN and XRD mineralogy on
		composites.
		Approximately 60 laboratory batch leach tests
		using a variety of reagent suites to assess
		potential extraction.
	The basis for assumptions or predictions regarding metallurgical amenability. It is	Metallurgical test work for Resource reporting on sulphide material conducted includes:
Metallurgical factors or assumptions	always necessary as part of the process of determining reasonable prospects for eventual economic extraction to consider potential metallurgical methods, but the assumptions regarding metallurgical treatment processes and	Detailed QEMSCAN and XRD mineralogy on 18 composites and a further 4 sets of mineralogy of flotation test products. Comminution testing includes 17 SMC SAG milling tests plus Ball Mill Work Indices.
	parameters made when reporting Mineral	Flotation testwork on a suite of six ore type
	Resources may not always be rigorous.	composites and four mining composites
	Where this is the case, this should be	comprising over 200 individual tests, over 20
	reported with an explanation of the basis of the metallurgical assumptions made.	locked cycle tests (LCT). LCT results were used as a basis for estimating
		metallurgical recovery.
		Recovery of intermediate products (enriched
		Cu/PGE concentrate and Ni/Co MHP) from concentrate enrichment of low grade nickel concentrates was estimated using pilot plant
		data from similar projects and scouting tests on
		samples from Gonneville.

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Criteria	JORC Code explanation Commentary
	The base case assumption for constraining the Resource is for flotation to produce a copper concentrate for sale, and a bulk nickel concentrate for enrichment in a midstream facility. Palladium recovery was predominantly into the copper concentrate. Cobalt is mineralogically associated with nickel and can be assumed to behave in a similar manner. Metallurgical recoveries used in the Resource pit optimisation were based on testwork completed to early 2024. Recovery algorithms calculated for each element were used as inputs into the pit optimisation and NSR calculations.
Environmental factors or assumptions	Assumptions made regarding possible waste and process residue disposal options. It is always necessary as part of the process of determining reasonable prospects for eventual economic extraction to consider the potential environmental impacts of the mining and processing operation. While at this stage the determination of potential environmental impacts, particularly for a greenfields project, may not always be well advanced, the status of early consideration of these potential environmental impacts should be reported. Where these aspects have not been considered this should be reported with an explanation of the environmental assumptions made. The environmental approval process has commenced however environmental considerations for potential mining have not yet been evaluated in detail. At this stage Chalice is unaware of any specific environmental issues that would preclude potential eventual economic extraction, subject to government approvals.
	Sample density determinations were carried out using the water displacement method. Incompetent oxide core samples from the
	Whether assumed or determined. If weathering profile are wax-coated prior to
	assumed, the basis for the assumptions. If density determination.
	determined, the method used, whether Density standards are employed in the density
	wet or dry, the frequency of the measurements, the nature, size and representativeness of the samples. determination process. Sample density determinations were carried out on all fresh rock core samples, and
	representative oxide samples resulting in ~80% of total drilled diamond core intervals having had density determinations completed.
Bulk density	The bulk density for bulk material must have been measured by methods that adequately account for void spaces (vugs, porosity, etc), moisture and differences between rock and alteration zones within the deposit. Incompetent oxide core samples are wax- coated prior to density determination.

	Sample density determinations were used to assign a bulk density value to the block model
	Discuss assumptions for bulk density estimates used in the evaluation process of the different materials. using a combination of assignment by geostatistical domain, and spatial estimation from density determinations from de-surveyed drillholes. Model tonnages are subsequently estimated on a dry basis.

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Criteria	JORC Code explanation	Commentary
		The Resource has been classified following due
		consideration of all criteria contained in Section
		1, Section 2 and Section 3 of JORC Code 2012
		Table 1 of ASX Announcement 23 April 2024. The
		Resource has been classified as either
		Measured, Indicated or Inferred based on data
		quality, sample spacing, mineralisation
		continuity, confidence in the geological
		interpretations, quality of the grade estimations
		and metallurgical processing knowledge.
		Primary mineralisation within the host Ultramafic
		intrusion has been classified as a combination
		of Measured, Indicated and Inferred.
		Measured, Indicated and Inferred wireframe
		volumes were developed from sectional
		interpretation strings, and model cells then
		coded with Resource Classification codes
		directly from the wireframe volumes.
		All fresh and transitional material within the
		Ultramafic intrusion, excluding the mostly barren
		dolerite, and informed by a reasonably
		consistent drill spacing of 80m, has been
		classified as Inferred, except in the northwest,
		where a number of new deeper holes are
		spaced wider than 80m, but are nevertheless
		deemed to be sufficient to infer geological and
		grade continuity at depth. Around the
		periphery of the drilling pattern, where
		extrapolation results in lower quality estimates,
Classification	The basis for the classification of the Mineral Resources into varying	Pd grade variography has informed a decision to limit the extrapolation of the Inferred material
	confidence categories.	to approximately 50m beyond the last drill hole.
		The 80m drill spacing corresponds to the
		nominal initial exploration drill hole spacing used
		for the deposit.
		An 80m drill spacing is considered by the
		Competent Person as being sufficient to imply,
		but not verify, geological and grade continuity
		for the deposit style.
		All fresh and transitional material within the
		Ultramafic intrusion, excluding the mostly barren
		granite, and dolerite dyke units, informed by a
		consistent drill spacing of 40m has been
		classified as Indicated. The selection of a 40m
		drill spacing distance for Indicated was based
		on results from a simulation-based drill hole
		spacing study carried out for the deposit
		indicating that the resource definition drill-out
		be conducted on a 40 m x 40 m drill spacing.
		Variogram ranges of the main economic grade
		variable, Pd, indicating that grade continuity
		does not exceed 40 m to 50 m within the high
		grade zones.
		Estimation quality metrics, such as slope of
		regression and average distance to sample
		were considered during the classification process. A 40 m drill spacing is considered by the
		Competent Person as being sufficient to allow
		estimationof the depositphysical

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Criteria	JORC Code explanation Commentary
	characteristics 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. All fresh and transitional material within the Ultramafic intrusion, excluding the mostly barren granite, and dolerite dyke units, informed by a consistent drill spacing of 10m has been classified as Measured. The selection of a 10m drill spacing distance for Measured was based on: Variogram ranges of the main economic grade variable, Pd, indicating that grade continuity averages 40m to 50m within the high Pd/sulphide zones and is on the order of hundreds of metres in the general Ultramafic zones. Estimation quality metrics, such as slope of regression and average distance to sample were considered during the classification process. A 10m drill spacing is considered by the Competent Persons as being sufficient to allow estimation of the deposit physical characteristics 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. All non-Ultramafic material (country rock and dykes) has not been classified and the Supergene unit has been considered ineligible to rise to level of the Measured category of confidence due to metallurgical uncertainty, hence it is capped at an Indicated classification where the drill spacing is 40m x 40m or tighter.
	Whether appropriate account has been taken of all relevant factors (i.e. relative confidence in tonnage/grade estimations, reliability of input data, confidence in continuity of geology and metal values, quality, quantity and distribution of the data). Appropriate account has been taken of all relevant criteria including data quality, sample spacing, mineralisation continuity, confidence in the geological interpretations, quality of the grade estimations and the availability of Modifying Factors.
	Whether the result appropriately reflects the Competent Person’s view of the deposit. The Mineral Resource appropriately reflects the Competent Person’s views of the deposit.
	Cube Consulting has undertaken internal peer
Audits or	The results of any audits or reviews of Mineral Resource estimates. reviews. Mr Mark Noppé of SRK Consulting completed in-stream reviews of previous Resource Estimates. No external review has
reviews
	been completed for this estimate.
Discussion of relative accuracy/ confidence	Where appropriate a statement of the relative accuracy and confidence level in the Mineral Resource estimate using an approach or procedure deemed appropriate by the Competent Person. For example, the application of statistical The Mineral Resource accuracy is communicated through the classification assigned to this Mineral Resource. The Resource has been classified in accordance with the JORC Code (2012 Edition) using a qualitative approach.

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Criteria	JORC Code explanation Commentary
	or geostatistical procedures to quantify the relative accuracy of the resource within stated confidence limits, or, if such an approach is not deemed appropriate, a qualitative discussion of the factors that could affect the relative accuracy and confidence of the estimate. All factors that have been considered have been adequately communicated in Section 1 and Section 3.
	The statement should specify whether it
	relates to global or local estimates, and, if local, state the relevant tonnages, which should be relevant to technical and economic evaluation. Documentation should include assumptions made and The Mineral Resource statement relates to a global tonnage and grade estimate. Grade estimates have been made for each block in the block model.
	the procedures used.
	These statements of relative accuracy and confidence of the estimate should be compared with production data, where available. No previous mining has taken place at the project, and production data is not available to reconcile against the block model estimates.

A-3-4 Section 4: Estimation and Reporting of Ore Reserves

Criteria	JORC Code explanation Commentary
Mineral Resource estimate for conversion to Ore Reserves	Description of the Mineral Resource estimate used as a basis for the conversion to an Ore Reserve. The Mineral Resource estimate used was prepared by Mike Millad and Mike Job of Cube Consulting and classified in accordance with the JORC 2012 guidelines. The basis of this Resource Estimate is as at 23 April 2024.
	Clear statement as to whether the Mineral Mineral Resources are inclusive of Ore
	Resources are reported additional to, or Reserves, noting that NSR cut-off calculations
	inclusive of, the Ore Reserves. differ between the two.
Site visits	Comment on any site visits undertaken by the Competent Person and the outcome of those visits. If no site visits have been undertaken indicate why this is the case. A site visit was conducted by the Competent Person on 27th June 2025 as part of the compilation of this Report.
	The Ore Reserve is reported as part of a Pre- Feasibility Study. The work undertaken to date has addressed all material Modifying
	The type and level of study undertaken to enable Mineral Resources to be converted to Ore Reserves. Factors required for the conversion of Mineral Resources to Ore Reserves and has shown that the mine plan is technically achievable and economically viable. The Ore Reserve
	has been based on parameters obtained
	from Chalice, from relevant technical studies and ongoing mining and processing parameters.
Study status
	The Code requires that a study to at least Pre-Feasibility Study level has been undertaken to convert Mineral Resources to Ore Reserves. Such studies will have been carried out and will have determined a mine plan that is technically achievable and economically viable, and that material Modifying Factors have been considered. Completed to PFS level.

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Criteria	JORC Code explanation Commentary
	Calculated value is based on a Net Smelter Return (NSR) to take account of the revenue from the palladium, platinum, copper, nickel, cobalt and gold allowing for metallurgical recoveries and payabilities for each and then offsets for royalties, transport and smelter deductions (penalty elements). A
Cut-off	The basis of the cut-off grade(s) or quality parameters applied. cut-off value of $17 has been applied to each oxide block for possible inclusion into Ore Reserves, whereas a cut-off of $23 has
parameters
	been applied to each sulphide block. A net value script was then applied to these blocks where a positive value was assigned as Ore Reserve status within the pit design. Net value was calculated as revenue less all operating costs. This mirrors the Whittle pit optimisation process.
Mining factors or assumptions	The method and assumptions used as reported in the Pre-Feasibility or Feasibility Study to convert the Mineral Resource to an Ore Reserve (i.e. either by application of appropriate factors by optimisation or by preliminary or detailed design). Open pit optimisations have been completed to generate a detailed mine design, schedule and cashflow model.
	The choice, nature and appropriateness of the selected mining method(s) and other mining parameters including associated design issues such as pre-strip, access, etc The Gonneville operation is proposed to use a conventional open cut excavator and truck mining fleet. This mining fleet is considered suitable for this type of surface mining operation.
	The assumptions made regarding geotechnical parameters (eg pit slopes, stope sizes, etc), grade control and pre- production drilling. Geotechnical analysis of the deposit was undertaken by Dempers and Seymour Pty Ltd. A geotechnical site investigation and laboratory testing was undertaken. Analysis of collected data, development of global significant geotechnical structural model & mining rock mass model and limit equilibrium & finite element stability analysis undertaken per geotechnical domain. Open pit comprises 7 geotechnical domains, overall slope angle per domain incorporates ramps and geotechnical berms, angles vary from 38-47 degrees for oxide (weathered) and 48- 54 degrees for fresh (sulphide). The proposed pit slopes are considered to be appropriate for the current pit design
	The Mineral Resource model has been provided by Cube Consulting and reviewed by Entech. The Mineral Resource Block Model
	The major assumptions made and Mineral was used for optimisation and mine planning
	Resource model used for pit and stope after inclusion of additional attributes to
	optimisation (if appropriate). become a Mining Model. A regularised Block
	Model was created with block sizes of 5 m x 5 m x 5 m which is considered suitable for the proposed mining method and equipment
	The mining dilution factors used. The regularisation process resulted in a mining dilution at $25 NSR of 4% and Ore loss of 4%.
	No additional mining dilution or ore loss
	The mining recovery factors used. factors have been applied after
	regularisation.

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Criteria	JORC Code explanation Commentary
	Any minimum mining widths used. A minimum mining width of 30m has been applied in the pit designs.
	The manner in which Inferred Mineral
	Resources are utilised in mining studies and No Inferred Mineral Resource is included in
	the sensitivity of the outcome to their the Ore Reserve.
	inclusion.
	The infrastructure requirements of the selected mining methods. The requirements to establish mine infrastructure were determined in the PFS and include: Mineral Processing Facility, Administration building and workshops, Tailings Storage Facility (TSF) Explosive Magazine, Landfill, Solar, BESS & diesel generation, Surface water management and onsite water storage dams, Transmission line & electrical substation, Water pipeline,
	The metallurgical process proposed and All metallurgical processes and recoveries are detailed in the Report, based on several
	the appropriateness of that process to the years of testwork, and the results from the
	style of mineralisation. work conducted have indicated that the
	processes are appropriate for the mineralisation present.
	Whether the metallurgical process is well- tested technology or novel in nature. Comminution and flotation processes are well-tested and proven technology for the feed samples. Leaching of Oxide and sulphide palladium is not currently practiced due to occurrence of PGMs almost universally being associated with base metal sulphides beneficiated by flotation. Gonneville has unique mineralogy where a significant portion of the palladium is in cyanide soluble form. Resin-in-pulp leaching is commonplace in gold operations.
Metallurgical
	The nature, amount and Metallurgical test work on oxide and fresh ore material conducted includes: Detailed TESCAN TIMA analysis was conducted on 163 samples and process stream during the PFS program. Comminution testing includes 95 SMC SAG milling tests plus >68 Ball Mill and abrasion
factors or
assumptions

	representativeness of metallurgical test work undertaken, the nature of the metallurgical domaining applied and the corresponding metallurgical recovery factors applied. work indices. HPGR and VRM pilot testing. Fresh ore testwork conducted on 7 development composites, 51 variability samples and 5 variability composites, comprising of over 800 individual tests and
	~29 locked cycle tests.
	Approximately 250 laboratory batch leach tests assessing 13 composites, 45 variability samples and a variety of conditions. All major adsorption, elution and regeneration steps have been designed based on testworkgeneratedfroma

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Criteria	JORC Code explanation Commentary
	synthetic solution grading 22 mg/L Ni, 17 mg/L Cu, 144 mg/L S and 0.11 mg/L Pd. The elution steps were also replicated on Au and Pt synthetic solution. Full flowsheet mass balances on the 7 fresh ore development composites, 1 oxide composite and 7 oxide variability samples were used as a basis for estimating metallurgical recovery and plant design. All 51 variability samples being progressively tested through the full flowsheet will be reported. The base case target for flotation is to produce separate copper (20% Cu target grade) and nickel (8% Ni target grade) concentrates for sale. All testwork has been carried out using Perth tap water, which is not expected to be substantially different to a treated Alkimos water supply.
	Any assumptions or allowances made for deleterious elements. No significant issues with deleterious elements have been identified throughout historical testing. MgO deportment to concentrate may incur a penalty at times. However, on average, the MgO grades have been lower than the required thresholds that trigger a penalty during the testwork program.
	The existence of any bulk sample or pilot A piloting campaign was conducted during
	scale test work and the degree to which such samples are considered representative of the orebody as a whole. the Scoping Study. A further pilot plant program is planned to commence as part of the Feasibility Study to confirm the PFS flowsheet using continuous, closed loop
	operation at bench scale.
	For minerals that are defined by a specification, has the ore reserve estimation been based on the appropriate mineralogy to meet the specifications? Metallurgical domaining is based on geochemical and geological logging data. This has been further refined with metallurgical testing, and the ability to produce separate saleable copper and nickel concentrates for fresh ores and leachable material from the oxide ore. inferring fresh and oxide material characteristics. Ore reserve has been defined by metallurgical models that define variable metal recovery to a target concentrate quality and leach residue grade based on chemical assays. Transitional ore recovery has been discounted as detailed in the Report.
	The status of studies of potential environmental impacts of the mining and processing operation. Details of waste rock characterisation and the consideration of Extensive work has been undertaken by Chalice to develop environmental baselines and define the programme of environmental surveys and studies required to support
Environmental	potential sites, status of design options formal environmental assessment during the study phase of the Project Formal referral of
	considered and, where applicable, the status of approvals for process residue storage and waste dumps should be reported. . the Project to State and Commonwealth governments was submitted in March 2024 which commenced the regulatory environmental approvals processes.

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Criteria	JORC Code explanation Commentary
	These baseline studies, along with other study deliverables such as waste characterisation and landform design (e.g. tailings storage facility, waste rock landform) inform the environmental impact assessment, targeting submission of the draft documents to regulators H2 2026.
Infrastructure	The existence of appropriate infrastructure: availability of land for plant development, power, water, transportation (particularly for bulk commodities), labour, accommodation; or the ease with which the infrastructure can be provided, or accessed. The project is located in proximity to the required infrastructure for mining being: Stage 1 requirement of 47MW serviced by ~27km transmission line to the existing SWIS Muchea Substation (MUC), Stage 2 requirement of 133WM provided by Clean Energy Link Chittering construction and connection to T9 (proximal to MUC). Stage 1 & 2 project water supplied from onsite contact water harvesting and ~63km water pipeline to Alkimos Waste Water Treatment Plant. Project has a Letter of Intent with Water Corporation for the supply of 8.4GL per year of recycled wastewater. The Great Northern Hwy (GNH), a primary distributor that provides connectivity to Perth, Fremantle Harbour, AMC Henderson, Port of Bunbury & Kwinana Bulk Terminal. Project access to GNH via ~28km to ~40km of regional distributor road network. The project is to be operated by a residential workforce that would reside in proximity to the project or Northern suburbs of Perth. A temporary construction camp is allowed for in the mine plan. The project has secured land in freehold for the development of mine infrastructure. Acquisition of select road reserves required in implementation.
	The derivation of, or assumptions made, regarding projected capital costs in the study. Processing capital costs have been provided by GR Engineering and NewPro. Mining capital costs are based on a first principles cost model. All other material capital costs are based on pricing received from third parties.
	The methodology used to estimate operating costs. Processing operating costs have been provided by GRES and NewPro. Mining operating costs are based on a first principles cost model assuming contractor mining. All other material operating costs are based on pricing received from third parties.
Costs

	A specification derived from concentrates
	Allowances made for the content of produced in testwork has been assessed by
	deleterious elements. copper and nickel smelters with negligible
	deleterious elements or penalties identified.
	The source of exchange rates used in the study. Cost models are in Australian dollars.

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	Derivation of transportation charges. Transportation charges have been provided by NMT.
	The basis for forecasting or source of treatment and refining charges, penalties for failure to meet specification, etc. Independent marketing advisor was engaged by Chalice to co-ordinate indicative terms from copper and nickel smelters. Four copper proposals and three nickel proposals were received and used as the basis for the offtake assumptions.
	The allowances made for royalties payable, both Government and private. Allowances for WA State Government royalties according to the Department of Treasury and Finance
Revenue factors	The derivation of, or assumptions made regarding revenue factors including head grade, metal or commodity price(s) exchange rates, transportation and treatment charges, penalties, net smelter returns, etc. A Net Smelter Return (NSR) was used for revenue. The NSR takes revenue from the palladium, platinum, copper, nickel, cobalt and gold allowing for metallurgical recoveries and payabilities for each and then offsets royalties, shipping and smelter charges/deductions. The following prices were applied to the optimisation process: Palladium - US$1,050 / oz Nickel - US$16,500 / t Copper - US$9,000 / t Platinum - US$1,000 / oz Gold - US$2,200 / oz Cobalt - US$30,000 / t AUD:USD – 0.65 The Competent Person considers this to be an appropriate commodity price assumption based on the current level of study and price environment at the time of the completion of the Ore Reserve work.
	The mine design and NSR prices for palladium, nickel and copper used reflect Chalice’s estimated ‘trough’ or low-point of future commodity price cycles, in real terms
	The derivation of assumptions made of (2025). Historical price trends and assessment
	metal or commodity price(s), for the of ‘resistance’ points in the industry cost
	principal metals, minerals and co-products. curves (typically 60-80th percentile) were
	used to guide the mine design prices. These prces also correspond to the low end of consensus estimated range from financial institutions.
Market assessment	The demand, supply and stock situation for the particular commodity, consumption trends and factors likely to affect supply and demand into the future. A customer and competitor analysis along with the identification of likely market windows for the product. For industrial minerals the customer specification, testing and acceptance requirements prior to a supply contract. Price and volume forecasts and the basis for these forecasts. World Platinum Investment Council (WPIC) identified in the September 2025 “Platinum Essentials” publication that Gonneville’s key commodity exposure, Palladium, has been in a deficit for the last 3 years. WPIC are forecasting a palladium deficit to continue for the next two years. The deficit is driven by a decrease in forecast mine production, noting ~90% of all mine supply is from South Africa and Russia, and an increase in forecast demand from autocatalyst consumption in hybrid and ICE vehicles, which account for ~85% of all Palladium demand consumption.

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Criteria	JORC Code explanation	Commentary
	For industrial minerals the customer	Gonneville will produce three saleable
	specification, testing and acceptance	products; copper-palladium-platinum-gold
	requirements prior to a supply contract.	(Cu-PGM) concentrate; nickel-cobalt-
		palladium-platinum (Ni-Co-PGM)
		concentrate; palladium-platinum-gold doré
		(PGM doré).
		The products are considered industry
		standard and commercially attractive to a
		broad range of potential customers. Global
		smelter complexes are the most likely
		customers for the Gonneville concentrates
		but downstream pCAM producers have also
		expressed interest in the Ni-Co-PGM
		concentrate.
		Chalice engaged an independent base
		metal marketing expert, Robert Aird, to
		advise and confirm the long term
		marketability and offtake terms for the
		Gonneville products.
		Engagement with potential offtake parties
		has confirmed the marketability of the
		Gonneville concentrates and dore, and
		potential offtake terms. Third parties have
		reviewed concentrate assays and
		conducted their own tests and provided
		indicative terms including payability and
		penalty ranges. Based on these discussions,
		no penalties are expected for deleterious
		elements.
		Concentrate volumes are based on the
		mining and processing production forecasts.
		At full scale production Gonneville will
		produce ~48ktpa of Cu-PGM concentrate
		and ~110ktpa of Ni-Co-PGM concentrate.
		Long term commodity price assumptions for
		coproducts are aligned with the 95th
		percentile of industry all-in sustaining cost
		curves.
		Revenue is determined by the metal content
		of the products, price assumptions with
		deductions for transport, shipping and
		refining costs.
		The costs of sales include the transport costs
		from mine to customer, and any commercial
		adjustments for non-revenue elements.
		The Gonneville Project is in the development
		phase, there are no off-take contracts
		currently in place.
		A discount rate of 8% per annum (real) has
		been applied, which is consistent with base
		metal projects and accounts for likely equity
Economic	The inputs to the economic analysis to produce the net present value (NPV) in the study, the source and confidence of these economic inputs including estimated inflation, discount rate, etc.	and debt financing costs. All material cost inputs are based on pricing received from third parties. Running the project financial model at the mine design / NSR inputs, the project
		generates a positive NPV:
		Palladium - US$1,050 / oz

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Criteria	JORC Code explanation Commentary
	Nickel - US$16,500 / t Copper - US$9,000 / t Platinum - US$1,000 / oz Gold - US$2,200 / oz Cobalt - US$30,000 / t AUD:USD – 0.65
	NPV ranges and sensitivity to variations in the significant assumptions and inputs. The optimisation shells upon which pit designs were based were generated at 99% of the base revenue. Project sensitivity analysis has been undertaken within the detailed financial model on key economic assumptions, Including commodity prices, Opex, Capex and FX. The Ore Reserve is most sensitive to commodity prices, in particular Palladium and Nickel prices where the Ore Reserve generates a positive NPV at the -5% and -9% range respectively.
	In August 2023 Chalice signed a Heads of Agreement (HoA) with the Shire of Toodyay for the establishment of a Community Fund that would operate once the Project reaches commercial production). Under the terms of the HOA, Chalice has agreed to provide funding for the delivery of community projects and programs that have been identified by the Shire of Toodyay and that align with Chalice’s eligibility criteria. The HOA will also form the basis for Chalice to
	The status of agreements with key establish similar community funds with other
Social	stakeholders and matters leading to social neighbouring Shires in the region.
	licence to operate. The Fund will aim to create lasting benefits for
	the local community, which are determined in consultation with the community. Chalice has actively and transparently engaged with local communities to keep people informed about the Project, to build relationships and better understand issues most relevant to the community. This engagement enhances our understanding of social licence, with the community expressing strong support for the future mine and its contribution to economic development
Other	To the extent relevant, the impact of the following on the project and/or on the estimation and classification of the Ore Reserves: Any identified material naturally occurring risks. No material naturally occurring risks have been identified.
	The status of material legal agreements No material risks have been identified.
	and marketing arrangements.
	The status of governmental agreements and approvals critical to the viability of the project, suchasmineral tenementstatus, The Project requires environmental approvals under the WA_Environmental Protection Act_ 1986 (_EP Act)_, andthe Commonwealth

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Criteria	JORC Code explanation Commentary
	and government and statutory approvals. There must be reasonable grounds to expect that all necessary Government approvals will be received within the timeframes anticipated in the Pre-Feasibility or Feasibility study. Highlight and discuss the materiality of any unresolved matter that is dependent on a third party on which extraction of the reserve is contingent. Environment Protection and Biodiversity Conservation Act 1999(EPBC Act). The Project was formally referred to the State and Commonwealth in March 2024 with both to be assessed with a Public Environmental Review. Studies are underway to support the environmental impact assessment anticipating submission of the draft documents H2 2026, with approvals targeted in H1 2028. Exploration Licences currently exist over the Chalice owned freehold titles for the Mine Development Area. Portions of this exploration tenure will be converted to a Mining Lease under the_WA Mining Act 1978_, aligning with the Mine Development Area. A Miscellaneous Licence under the_Mining Act_ _1978_will be sought for the power and water infrastructure.
	The Mineral Resources above an in-situ variable economic cut-off grade within the designed open pit has been modified by the
	The basis for the classification of the Ore Reserves into varying confidence categories. application of suitable modifying factors and has been classified as Probable, based on the Indicated classification of the Mineral Resource estimate. The level of work
	undertaken through pit optimisation studies and pit designing is considered sufficient for the classification of Probable Ore Reserves.
Classification
	Whether the result appropriately reflects the Competent Person’s view of the deposit. Mr. Daniel Donald, the Competent Person for this Ore Reserve estimation, has reviewed the work undertaken to date and considers that it is sufficiently detailed and relevant to each of the deposits to allow those Ore Reserves derived from Indicated Mineral Resources to be classified as Probable.

	The proportion of Probable Ore Reserves that have been derived from Measured Mineral Resources (if any). No Probable Ore Reserves have been based on Measured Mineral Resources
Audits or reviews	The results of any audits or reviews of Ore Reserve estimates. The Ore Reserve has been estimated by Independent consultants Entech Pty Ltd. Entech have undertaken internal peer reviews during the process.
	Where appropriate a statement of the relative accuracy and confidence level in the Ore Reserve estimate using an
	approach or procedure deemed The Competent Person deems that the
Discussion of	appropriate by the Competent Person. For example, the application of statistical or geostatistical procedures to quantify the methodology applied to arrive at the Ore Reserve estimate is appropriate.
relative
accuracy/	relative accuracy of the reserve within stated confidence limits, or, if such an The overall accuracy of the cost estimate used in the estimation of these Ore Reserves
confidence
	approach is not deemed appropriate, a is ±25%.
	qualitative discussion of the factors which could affect the relative accuracy and confidence of the estimate.

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Criteria	JORC Code explanation Commentary
	The statement should specify whether it relates to global or local estimates, and, if local, state the relevant tonnages, which should be relevant to technical and economic evaluation. Documentation should include assumptions made and the procedures used. The statement relates to global estimates of a mine scale.
	Confidence in the application of the modifying factors is appropriate for the estimate.
	Accuracy and confidence discussions A degree of uncertainty is associated with
	should extend to specific discussions of any geological estimates and the Ore Reserve
	applied Modifying Factors that may have a classification reflects the level of confidence
	material impact on Ore Reserve viability, or in the Mineral Resource.
	for which there are remaining areas of There is a degree of uncertainty regarding
	uncertainty at the current study stage. estimates of mining modifying factors,
	It is recognised that this may not be geotechnical and processing parameters
	possible or appropriate in all that are of a confidence level reflected in
	circumstances. These statements of relative the level of the study.
	accuracy and confidence of the estimate There is a degree of uncertainty in the
	should be compared with production commodity price used however the
	data, where available. Competent person(s) are satisfied that the
	assumptions used to determine the economic viability of the Ore Reserve are based on reasonable current data.

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AI assistant

CHALICE MINING LIMITED — Capital/Financing Update 2025

64649_rns_2025-12-07_2eea1cd6-cc6f-47b5-aff7-528371c8323e.pdf

ASX Announcement

8 December 2025

Gonneville Palladium-Nickel-Copper Project PFS

Highlights

Investor Conference Call

Project overview

Key study outcomes and metrics

Production target

Financial return metrics

Detailed metrics by stage

Cost profile

Industry competitiveness

Sensitivity analysis

Development plan overview

Mining

Process plant

Product marketing and offtake

Supporting infrastructure and workforce

Regulatory approvals

Community

Environment

Project funding

Development timeline and next steps

For further information, please visit www.chalicemining.com, or contact:

Corporate Enquiries

Media Enquiries

Follow our communications

Pre-Feasibility Study

Table of Contents

1. Project location and history

1.1 Mine Development Area

1.2 Landform and topography

1.3 Flora and fauna

1.4 Surface water and ground water

1.5 Climate

2. Development plan overview

2.1 Mining

2.2 Process plant

2.3 Product marketing and offtake

2.4 Supporting infrastructure and workforce

3. Geology and Mineral Resources

3.1 Geology

3.2 Mineralisation

3.3 Mineral Resources

4. Geotechnical

4.1 Data collection

4.2 Rock Mass Strength

4.3 Output data

4.4 Pit slope design

4.5 Stability analysis

5. Mining and Ore Reserves

5.1 Dilution analysis

5.2 Open-pit mine design

5.3 Open-pit mine schedule

5.4 Ore Reserve

6. Metallurgy and processing

6.1 Metallurgical testwork

6.1.1 Feed samples

6.1.2 Process water samples

6.2 Mineralogy

6.2.1 Sulphide

6.2.2 Oxide

6.3 Process and flowsheet design

6.3.1 Trade-off studies

6.3.2 Oxide comminution

6.3.3 Sulphide comminution

6.3.4 Sulphide flotation concentration

6.3.5 Leach

6.4 Geo-met and recovery modelling

6.4.1 Oxide recovery

6.4.2 Transitional recovery

6.4.3 Fresh sulphide recovery

7. Tailings and waste management

7.1 Tailings Storage Facility design

8. Infrastructure and logistics

8.1 Process water supply

8.2 Power supply

CHALICE MINING LIMITED Capital/Financing Update 2025