5 August 2026

Central Asia Metals PLC

(the 'Group', the 'Company' or 'CAML')

Mineral Resource and Ore Reserve Update

Central Asia Metals PLC (AIM: CAML) provides an update on its Mineral Resource and Ore Reserve Estimates.

As part of the proposed acquisition of Cygnus Metals Ltd (ASX: CY5, TSXV: CYG, OTCQB: CYGGF) ('Cygnus') via an Australian scheme of arrangement, as announced on 2 June 2026, CAML commissioned SLR Consulting (UK) Ltd ('SLR') to prepare Mineral Resource and Ore Reserve Reports in accordance with the guidelines of the Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves (2012 Edition) (the 'JORC Code') to disclose information relevant to investors about:

-      CAML's wholly-owned Kounrad in-situ dump-leach, solvent extraction-electrowinning (SX-EW) copper operation in Kazakhstan ('Kounrad'); and

-      CAML's wholly-owned Sasa zinc-lead mine in North Macedonia ('Sasa').

The reports include, amongst other information, Mineral Resource and Ore Reserve Estimates with effective dates of 31 December 2025. The reports are available on CAML's website (at www.centralasiametals.com/investors/reports-and-presentations). Summaries of the Mineral Resource and Ore Reserve Estimates for Sasa and Kounrad are set out below.

Furthermore, as part of the documentation required for the proposed acquisition of Cygnus, CAML also commissioned SLR to prepare Technical Reports about Kounrad and Sasa in accordance with the disclosure requirements of National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101). Those reports will be available on the Canadian Securities Administrators' website under Cygnus' corporate profile at www.sedarplus.ca (SEDAR+) concurrent with the planned publication of the Scheme Booklet to be despatched to Cygnus shareholders in relation to the Scheme.

Kounrad (100% owned)

Kounrad Mineral Resource Estimate

A summary of the audited Mineral Resource Estimate as of 31 December 2025, is shown in Table 1.

Table 1: Audited Mineral Resource Statement for Kounrad (effective date of 31 December 2025)

Notes

1.        Mineral Resources are reported in accordance with the guidelines of the JORC Code (2012).

2.        Cu grade is total copper.

3.        Mineral Resources are estimated at a cut-off grade of 0% Cu.

4.        Mineral Resources are estimated using a copper price of US$11,080 per tonne.

5.        Mineral Resources are estimated using a calculated leach recovery of 18.5%.

6.        Mineral Resources are estimated using average dry densities ranging from 1.875 to 2.07 t/m³.

7.        Kounrad is an in-situ dump-leach operation, and therefore no minimum mining width has been applied.

8.        Mineral Resources are inclusive of Ore Reserves.

9.        Mineral Resources that are not Ore Reserves do not have demonstrated economic viability based on a prefeasibility study or feasibility study.

10.     Contained metal refers to estimated contained metal in the ground not adjusted for metallurgical recovery.

11.     Totals may not represent the sum of the parts owing to rounding.

Geology and geological interpretation

The Kounrad open pit is located adjacent to the waste dumps and comprises a porphyry copper deposit that was mined from 1936 to 2005. Copper sulphide ore was selectively mined, while material classified as waste and uneconomically treatable materials at the time were dumped and formed the existing Kounrad waste dumps. The waste material is comprised of oxide, mixed and sulphide granodiorites at grades of around 0.1% Cu. The Eastern Dumps contain mainly oxide material whereas the larger Western Dumps contain sulphide and mixed material. The smaller Northern Dumps also contain mixed material.

Malachite, azurite, chrysocolla and chalcopyrite (partly replaced by hematite and pyrite) are typical ore minerals of the oxide dumps. The sulphide dumps are characterised by vein and veinlet-disseminated sulphide mineralisation with covellite, chalcopyrite, bornite and pyrite. Long-term exposure of the dumps has triggered natural bacterial oxidation, converting a significant proportion of the refractory sulphide material into acid-soluble forms which are more amenable to leaching.

The Kounrad dumps are located on weathered bedrock of which the upper 10 to 40 metres consists of clay material that forms a low permeability layer. This is a critical factor in capturing the leach solutions after percolating through the dumps.

Sampling and sub-sampling techniques

Sampling of the Kounrad copper dumps included trenching, trial pits and reverse circulation (RC) drilling, and is summarised below:

-      Pre-2007: an historical programme of channel sampling was undertaken on the Eastern Dumps and included the collection of 2,409 samples of 0.5 metre lengths taken from the surface of the dumps.

-      2007: exploration work consisted of 21 RC drillholes in Dumps 6, 7, and 9-10. Samples were assayed for total copper (Cu_total) and acid soluble copper (Cu_acid).

-      2008-09: exploration work comprised 85 RC drillholes and 10 channel trenches with samples assayed for Cu_total and Cu_acid.

-      2010: exploration work included 137 trial pits, 9 surface trenches and a further 13 pits excavated for metallurgical sampling. Samples were assayed for Cu_total, Cu_acid.

-      2011: RC drillholes were drilled in Western Dumps, 1, 15 and 16, and Eastern Dumps 5, 6, 7, 9 and 10. A total of 98 holes were carried out and drilled through the full thickness of the dumps. Samples were assayed for Cu_total and Cu_acid.

-      2012: 131 RC drillholes were completed in Dumps 2, 13, 20, 15, 16, 20, 21 and 22 and the Northern Dumps.

Sampling of RC drilling was undertaken at 3 metre intervals. Samples were initially sub-sampled at the drill site to provide a 5-kilogramme sample that was then split to provide a sub-sample that was dispatched to the VNIITSvetmet laboratory in Ust-Kamenogorsk for sample preparation. The samples were crushed in a jaw crusher to <2 millimetres followed by secondary crushing to <1 millimetres. The sample was then split to provide <0.5 kg which was pulverised to 74 microns and further split to provide a 250 gramme sample for analysis.

Drilling techniques

All drilling at the Kounrad dumps has been undertaken using RC drilling. The number of drillholes completed by each drilling programme is provided below:

-      2007: 21 RC drillholes to depths of 2.5 metres.

-      2008-09: 85 RC drillholes (1,971 metres) to depths of 8-30 metres.

-      2011: 98 RC drillholes (3,213 metres) through the full thickness of the dumps.

-      2012: 131 RC drillholes (4,107 metres) through the full thickness of the dumps.

Limited information exists regarding the pre-2011 drilling. The 2011 and 2012 drilling was completed by the drilling contractor AK Niyet Burga using Nemek 814 BE and HYDCO - 300 trailer drill rigs, with a hole diameter of 125 millimetres.

Classification

The Kounrad waste dumps have been classified as wholly Indicated Mineral Resources. Drilling of the dumps has been undertaken on a maximum spacing of 200-100 metres. The composition of the dump material, the copper grades and amenability to leaching is well understood through exploration, test work and production.

Sample analysis and method

Most of the assaying was conducted at the VNIITSvetmet laboratory in Ust-Kamenogorsk. Pit and trench samples from 2010 were assayed at CenterGeoAnalyt in Karaganda. The Alex Stewart laboratory in Moscow and CenterGeoAnalyt were used for check analysis.

At the time of analysis, the VNIITSvetmet laboratory held several accreditations, including ISO/IEEC 17025-2007 (accreditation certificate number KZ.И. 07.0480 dated 27 August 2009) and ISO 9001:2008. The accreditation of the CenterGeoAnalyt and Alex Stewart laboratories at the time of the analysis is not known.

Total copper was analysed using aqua regia acid digestion with ICP-MS finish. Analysis for acid soluble copper was carried out on a 50 gramme sample treated with a 5% solution of sulphuric acid and heated for 30 minutes. The leach residue was washed with water, and the sample analysed by ICP. The difference in the total copper and copper in the leach residue was calculated as the acid-soluble copper content. A summary of the sample analysis and method is provided below:

-      Pre-2007: Limited information exists regarding the assaying methods used for the historical channel sampling of the Eastern Dumps.

-      2007-09: Samples were assayed at the VNIITSvetmet laboratory in Ust-Kamenogorsk. No information is available of the QA/QC procedures used for the 2007-09 exploration works.

-      2010: Samples were sent for preparation and assay at the CenterGeoAnalyt laboratory in Karaganda. VNIITSvetmet laboratory was used for external check analysis. Analysis was undertaken for Cu_total and Cu_acid. A total of 30 external pulp duplicates and 59 internal control samples were analysed.

-      2011: Samples were assayed at the VNIITSvetmet laboratory. The Alex Stewart laboratory was used for external check analysis. Analysis was undertaken for Cu_total and Cu_acid. A total of 918 samples were assayed, of these 30 (3%) were selected as internal laboratory duplicates and 75 (8%) were selected as umpire duplicates for analysis at Alex Stewart. Each batch of 60 samples included 1 barren (blank) sample, which consisted of blank granite material. A total of 16 blank samples were assayed.

-      2012: Samples were assayed at the VNIITSvetmet laboratory. Alex Stewart laboratory was used for external check analysis. Analysis was undertaken for Cu_total and Cu_acid. A total of 1,364 samples were assayed at the VNIITSvetmet laboratory, of these 114 (8%) were selected as internal laboratory duplicates and 137 (10%) were selected as umpire duplicates for analysis at Alex Stewart. Each batch of 60 samples included 1 barren (blank) sample, which comprised blank granite material. A total of 20 blank samples were assayed.

Limited information exists regarding the procedures used for the pre-2007 sampling and 2007 drilling programmes, however, these were undertaken in the Eastern Dumps which have now been extensively leached. Limited information exists regarding the procedures used for the 2008-09 drilling programme. A statistical analysis comparing the results of this programme with the 2011-12 drilling identified no significant bias in the copper grades to be present that would materially affect the Mineral Resource Estimate. The 2010-12 programmes were undertaken using appropriate sample preparation and analyses procedures and no significant issues were identified in the assaying of the QC samples.

Estimation methodology

The Mineral Resource Estimates for the Kounrad in-situ dump-leach operation were produced by Sary Kazna LLP (SK) using a polygonal based approach and were audited by SLR.

An initial Mineral Resource Estimate for Dumps 6, 7 and 9-10 had an effective date of 1 December 2011, and included the 2011 drilling (no drilling was undertaken on these dumps in 2012). This was before leaching operations commenced on these dumps in April 2012. An initial Mineral Resource Estimate for the remaining dumps had an effective date of 17 May 2013, and included the 2012 drilling. No further exploration has been undertaken since this time. The initial Mineral Resource Estimates are updated annually by SK to account for depletion from on-going leaching operations.

To audit the SK Mineral Resource Estimates, SLR undertook a check estimate using a 3D block modelling approach using Leapfrog, Supervisor and Datamine software. The sample database included: 1) Pre-2007 channel sampling; 2) 2007 drilling; 3) 2008-09 drilling; 2010 trial pit sampling, metallurgical trial pit sampling and trenching; 4) 2011 drilling; 5) 2012 drilling.

Wireframes of the topographic surface of the dumps were constructed based on surveys carried out by SK surveyors using total station equipment. A surface defining the base of the dumps was constructed by SLR using the drillhole logging of basement material identified in the 2011 and 2012 campaigns.

The sample database and the volumetric block model were coded based on the dump IDs. High-grade Cu assays were top cut where required. A 3.0 metre composite length was used. Variogram contour maps and variography was attempted; however, robust variograms could not be produced to support a kriged estimate.

A block model was constructed using a parent block size of 50 x 50 x 3 metres and sub cell splitting was enabled to more accurately represent the dump volumes. Average densities were based on density test work from sampling of trial pits in 2012 and were applied on a dump-by-dump basis. Grade estimation was undertaken for copper and acid soluble copper. Inverse distance weighting (IDW2) was used as the principal estimation method. Domain boundaries of the dumps were treated as hard boundaries. Grade estimation was undertaken into the parent blocks. Grade estimation used a three-pass plan. The first search size was based on the 200 x 100 metre sample spacing. The second and third searches used x2 and x3 expansions to estimate any remaining blocks. First-pass block estimates were required to be informed by a minimum of 4 composites and a maximum of 12 composites, with a maximum of 2 composites per drillhole. Sample requirements were relaxed in later estimation passes to ensure that all blocks received estimated grades. The channel sample data were restricted to estimate only those blocks located near the surface to prevent over extrapolation of these data at depth. Directional control strings were used to orientate the search ellipses based on the dump profiles.

The block model was validated by 1) a visual comparison of composites and block grades; 2) a global statistical comparison of the composites and block grades by domain; 3) swath plots.

A comparison between the SLR check estimate and the SK Mineral Resource Estimate identified no material differences. Globally, the models reported similar grades, total reported tonnes were within 1%, and the total contained metal was within 2.5%. SLR accepts the SK Mineral Resource Estimate and the reported Mineral Resources are based on the SK estimate.

The copper recovered from the dumps since the start of leaching operations in April 2012 is estimated by SK metallurgical staff based on the results of laboratory and pilot plant testing, assumptions of leach and solution migration rates and cathode production information. Overall, a total of 179.3 kilotonnes of copper has been recovered and 75.4 kilotonnes of copper remains to be recovered. The Mineral Resources were depleted to account for the copper that has been recovered by the operation since April 2012.

Mineral Resources were reported inside of the concession boundary. The small Dump 3 (approximately 500,000 cubic metres) was excluded from the Mineral Resource Estimate.

Cut-off grade

Kounrad is an in-situ dump-leach operation and is not selectively mined. A cut-off grade of 0% Cu is therefore used to report the Mineral Resources.

Mining, metallurgical and other material modifying factors

Kounrad is an in-situ dump-leach operation and no mining is undertaken.

The metallurgical process is well-tested and has been operating since 2012. Leach recoveries attained by the operation are estimated at 51% for the Eastern (oxide) Dumps, while the Western Dumps, being more refractory, respond less favourably to leaching and recoveries ranging from 35% (sulphide material) to 42% (mixed material) have been achieved.

The stated Mineral Resources are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues. There is no known mining, metallurgical, infrastructure or other factor that materially affects this Mineral Resource Estimate at this time.

Comparison with previous Mineral Resource Estimate

A Mineral Resource Estimate was prepared by SK before the start of leaching operations at Kounrad. A comparison of the previous and the current Mineral Resource Estimates is shown in Table 2. The estimates are in accordance with the guidelines of the JORC Code (2012).

The total Mineral Resource tonnage at Kounrad remains the same because no material has been mined or removed. A reduction in contained copper metal from 587.3 to 407.9 kilotonnes and a reduction in copper grade from 0.10% to 0.07% Cu is due to the recovery of copper by the leaching operations from April 2012 to 31 December 2025.

Table 2: Comparison of previous and current Mineral Resource Estimates

Notes

1.        The effective date of the Mineral Resource Estimate for Dumps 6, 7 and 9-10 is 1 December 2011.

2.        The effective date of the Mineral Resource Estimate for Dumps 2, 5, 1, 1a, 13, 15, 16, 20, 21, 21a, 22 and Northern Dumps 1 & 2 is 17 May 2013.

3.        Cu grade is total copper.

4.        Mineral Resources are estimated at a cut-off grade of 0% Cu.

5.        Contained metal refers to estimated contained metal in the ground not adjusted for metallurgical recovery.

6.        Totals may not represent the sum of the parts owing to rounding.

Kounrad Ore Reserve Estimate

The audited Ore Reserve statement is shown in Table 3.

Table 3: Audited Ore Reserve Statement for Kounrad (effective date of 31 December 2025)

Notes

1.        Ore Reserves are reported in accordance with the guidelines of the JORC Code (2012).

2.        Cu grade is total copper.

3.        Ore Reserves are estimated at a cut-off grade of 0% Cu.

4.        Ore Reserves are estimated using a copper price of US$11,080 per tonne.

5.        Ore Reserves are estimated using a calculated leach recovery of 18.5%.

6.        Kounrad is an in-situ dump-leach operation, and therefore no dilution or mining recovery factors have been applied.

7.        Contained metal refers to estimated contained metal in the ground not adjusted for metallurgical recovery.

8.        Totals may not represent the sum of the parts owing to rounding.

Material Assumptions

Kounrad is an operating in-situ dump-leach operation focussed on the recovery of copper from waste dumps adjacent to the historical Kounrad open pit. The leaching operation has been in continuous production since April 2012 and costs and revenue are well understood. Modifying factors were based on actual mineral processing, infrastructure, environmental and social, market and economic model information.

A financial model for the LoM showed the Ore Reserves to be economic based on the assumptions used and a copper price of US$11,080 per tonne.

Classification

The Ore Reserve Estimate is based on the Mineral Resource model. Indicated Mineral Resources were converted to Probable Ore Reserves by applying modifying factors. No Proved Ore Reserves were classified because no Measured Mineral Resources were estimated.

No Inferred Mineral Resources were estimated and no Inferred Resources were included in the Ore Reserve Estimate or the economic analysis.

Mining method and other mining assumptions

The operation is an in-situ dump-leach and no mining is required. Dilution and mining recovery factors are not applicable and no minimum mining widths have been applied.

Processing method and other processing assumptions

The dumps are leached in situ with raffinate solution at pH 1.2 to dissolve the copper mineralisation. The solution percolates through the dumps to a layer of low permeability clay at the base of the dumps. The pregnant leach solution (PLS) is collected at the base of the dumps by an HDPE lined interceptor trench and pumped to a series of holding ponds. From these ponds, the PLS is treated by the SX-EW plant, located south of the Eastern Dumps. The east plant site includes the SX-EW plant, solution ponds, boiler houses, stores and offices. PLS solutions from the Western Dumps are pumped a distance of 12.5 kilometres to the east plant for copper recovery.

A leach recovery of 51% has been achieved for Eastern Dumps 6, 7, 9-10. The Western Dumps, being more refractory, respond less favourably to leaching with recoveries ranging from 35% (sulphide material) to 42% (mixed material) being achieved.

The metallurgical process is well-tested and has been operating since 2012. A total of 178,961 tonnes of copper cathode has been produced since the start of the operation in April 2012 up to 31 December 2025. In future years, production is planned to decline as fewer dumps remain to be leached and those that remain are lower grade and comprise mainly sulphide or mixed material. The production schedule contains a total of 74,459 t of copper cathode that will be produced from 2026-34.

Based on the Ore Reserve Estimate and the estimated metallurgical recoveries there is sufficient leachable metal to support the production plan.

The copper cathode produced is 99.99% Cu. No significant levels of deleterious elements are present.

Cut-off grade

Kounrad is an in-situ dump-leach operation and is not selectively mined. A cut-off grade of 0% Cu is therefore used to report the Ore Reserves.

Estimation methodology

The Mineral Resource Estimate on which the Ore Reserve Estimate has been based was prepared by SK and dated 31 December 2025.

Indicated Mineral Resources were used as the basis of the Ore Reserve Estimate and were converted to Probable Ore Reserves based on the demonstrated technical and economic viability of the production schedule.

Material modifying factors

Permits held by SK and Kounrad Copper Company LLP (KCC) for the Property are sufficient to ensure that processing activities are conducted within the regulatory framework. SK and KCC are wholly owned subsidiaries of CAML.

SK holds a concession area of approximately 14.8 square kilometres. The concession relates to the exploration and extraction of copper from the waste dumps and has an expiry of 20 August 2034. The east plant and solar farm are outside of the concession and operate under lease arrangements. Dumps located outside of the concession are not included in the Mineral Resource or Ore Reserve Estimates.

Land tenure for the Kounrad operation is held under a state long-term lease arrangement. An alignment between the concession and the land tenure is required to ensure full consistency between these boundaries and is a planned transfer.

All major infrastructure required by the operation is in place and no additional significant infrastructure projects are currently planned.

The stated Ore Reserves are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues. There is no known mining, metallurgical, infrastructure or other factor that materially affects this Ore Reserve Estimate at this time.

Comparison with previous Ore Reserve Estimate

There are no previous Ore Reserve Estimates reported for the Kounrad in-situ dump-leach operation.

Production target

As at 31 December 2025 there remained approximately 75,400 tonnes of copper that could be extracted from Kounrad, which should ensure a life of operation until expiry of the current licence in 2034. The production schedule contains a total of 74,459 tonnes of copper cathode that will be produced from 2026-34. Based on the Ore Reserve Estimate and the estimated metallurgical recoveries there is sufficient leachable metal to support the production plan.

Notwithstanding the production schedule shown below, CAML's copper production guidance for 2026 remains unchanged at 12,000 to 13,000 tonnes, with the figure shown in the schedule representing a minimum.

Table 4: Summary of Kounrad production schedule

Notes

Reserves and Resources underpinning the production target

The production target is solely underpinned by Ore Reserves.

The estimated Ore Reserves underpinning the production target have been prepared by a Competent Person in accordance with the requirements of the JORC Code.

Sasa (100% owned)

Sasa Mineral Resource Estimate

A summary of the audited Mineral Resource Estimate as of 31 December 2025 is shown in Table 5.

Table 5: Audited Mineral Resource Statement for Sasa (effective date of 31 December 2025)

Notes

1.        Mineral Resources are reported in accordance with the guidelines of the JORC Code (2012).

2.        Mineral Resources are estimated using NSR cut-off values of US$53 per tonne for sub-level caving, US$65 per tonne for cut and fill stoping and $60 per tonne for long-hole stoping.

3.        Mineral Resources are estimated using metal prices of US$3,041 per tonne for zinc, US$2,506 per tonne for lead and US$31 per ounce for silver.

4.        Mineral Resources are estimated using metallurgical recoveries of 94% for lead, 82% for zinc and 75% for silver.

5.        Mineral Resources are estimated using average dry densities ranging from 2.7 to 4.0 t/m³.

6.        A minimum thickness of 1.5 metres was used to model the mineralised zones.

7.        Mineral Resources are inclusive of Ore Reserves.

8.        Mineral Resources that are not Ore Reserves do not have demonstrated economic viability based on a prefeasibility study or feasibility study.

9.        Contained metal refers to estimated contained metal in the ground not adjusted for metallurgical recovery.

10.     Totals may not represent the sum of the parts owing to rounding

Geology and geological interpretation

The Sasa deposits of Svinja Reka and Golema Reka are located in the Serbo-Macedonian massif which comprises greenschist and amphibolite facies metamorphic rocks, Precambrian to Palaeozoic in age, which have been variably intruded by andesitic to trachytic volcanic rocks during the Tertiary. Lead-zinc-silver mineralisation at Sasa occurs as bedding concordant deposits hosted predominantly by quartz-graphite schist and marbles of Lower Palaeozoic age at Svinja Reka, while at Golema Reka it is primarily developed within gneiss host rocks.

High-temperature hydrothermal fluids and bedding-parallel faulting (related to the intrusion of Tertiary volcanics) is responsible for metasomatism of the host sediments to develop skarn and base metal mineralisation. The well-defined, partially exploited lenses of lead-zinc-silver mineralisation dip at approximately 35 degrees to the southwest and typically range in true thickness from between 2 and 30 metres.

The Sasa deposits are considered to be polymetallic lead-zinc-silver skarn-hydrothermal replacement systems developed within favourable metamorphic host rocks and influenced by later structural reactivation.

The Kozja Reka occurrence is located northwest of Golema Reka and was historically mined, however has insufficient exploration to define Mineral Resources.

Sampling and sub-sampling techniques

Surface and underground diamond core drilling is used to provide samples as the basis for the Mineral Resource Estimates. Samples from drill core are collected by Sasa geological staff. A sample interval of 0.3 to 1.0 metres is used. Logging, sampling and sample cutting (half core) are undertaken at the on-site core logging facility at Sasa. Sample preparation is undertaken at the Sasa analytical laboratory where half core is crushed to -3 millimetres and then dried in an oven at 130 degrees centigrade. The sample is passed through a riffle splitter to derive a 50% split, which is pulverised using a disc mill to give a -0.74 micrometres, 50 gramme pulp sample for analysis.

Drilling Techniques

Underground core drilling comprises most of the drilling completed at Sasa. Underground drillholes are completed using 76-millimetre (NQ) diameter, 36-millimetre diameter (BQ) or AX diamond core in up-holes.

Surface core drilling is less frequently undertaken and uses 76-millimetre diameter, HQ or NQ diamond core.

Classification

No Measured Mineral Resources were classified owing to the spatial complexity and local variability of the mineralisation. Indicated Mineral Resources were classified based on a minimum of four drillholes and a spacing of 50 metres or less. Inferred Mineral Resources were classified based on a minimum of two drillholes and a spacing of up to 150 metres.

Sample analysis and method

The Sasa laboratory analyses the 50 gramme pulp samples for lead, zinc and silver by atomic absorption (AA) using three-acid digestion. The Sasa laboratory is not currently accredited for the analytical method used for drill samples. A systematic Quality Assurance/Quality Control (QA/QC) system is used to monitor the accuracy and precision of assaying. Routine QA/QC samples (certified standards, blanks and coarse reject duplicates are submitted in the sample stream by Sasa geological staff. The results of the internal QA/QC analysis are considered acceptable.

In 2021 and 2024, the Sasa laboratory submitted check samples to the accredited Eurotest-Control EAD laboratory in Sofia, Bulgaria. The results of the check analysis for both periods are considered acceptable and identified no significant issues with the Sasa laboratory analysis.

Estimation methodology

Underground mapping combined with diamond drilling information has been used to interpret the mineralised zones and define three-dimensional solids in Leapfrog Geo software. A nominal 2% Pb+Zn cut-off plus lithology was used to constrain the mineralised domains and improve the continuation of the orebodies, although, since the contacts are generally sharp, some weaker mineralisation was included in the model.

Drillhole samples located within the domains were selected and coded by the principal domains. The samples were composited to 1 metre. Grade capping was applied to the composites based on appropriate population breaks indicated in the log probability plots. Geostatistical analyses and variography was undertaken. A block model was generated within the domain wireframes using parent blocks of dimensions 3.5 metres east by 14 metres north and 7 metres Relative Level.

Sub-cell splitting was applied to better represent the natural geometry and distribution of the mineralisation and reflect the ability to discriminate between higher grade and lower grade blocks or waste blocks at a size corresponding to the proposed mining methods. Grades were estimated into the block model using Ordinary Kriging for lead, zinc and silver. Multiple search passes were used to ensure all blocks received estimated grades. The domain boundaries were considered as hard boundaries. Bulk density values were interpolated using Nearest Neighbour and/or a regression formula based on estimated lead grades. The block model was visually and statistically validated by comparing the estimated block grades relative to the original sample results.

Cut-off grade

Net smelter return (NSR) was calculated to determine the value of each individual stope and used metal prices, payable metals considering lead, zinc and silver grades, metallurgical recoveries, and realisation costs. Metal prices used in the calculation of NSR for Mineral Resources were US$3,041 per tonne for zinc, US$2,506 per tonne for lead, and US$31 per ounce for silver. Metallurgical recoveries used in the calculation of NSR were 94% for lead, 82% for zinc and 75% for silver.

The NSR calculation also considered the planned underground mining methods based on orebody geometry and mining level. NSR cut-off values (COVs) of US$53 per tonne for sub-level caving (SLC), US$65 per tonne for cut and fill stoping (C&F) and US$60 per tonne for long-hole stoping (LHS) were used to estimate the Mineral Resources.

Mining, metallurgical and other material modifying factors

Mining of the Svinja Reka deposit is planned to continue using SLC, C&F and LHS stoping methods. Mining of the Golema Reka deposit is assumed to use LHS for the purposes of resource estimation.

The Sasa processing plant has been operating since 2006 and comprises a three-stage crushing circuit and conventional grinding, classification and selective flotation circuits to produce lead and zinc concentrates that are thickened and filtered for shipment by truck. Actual metallurgical recoveries achieved by the processing plant are 94% for lead, 86% for zinc and 75% for silver.

The stated Mineral Resources are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues. There is no known mining, metallurgical, infrastructure or other factor that materially affects this Mineral Resource Estimate at this time.

Comparison with previous Mineral Resource Estimate

A previous Mineral Resource Estimate for Sasa was prepared by CAML and reported in accordance with the JORC Code (2012) and with an effective date of 31 December 2024. A comparison of the current and the previous Mineral Resource Estimates is shown in Table 6.

The total Svinja Reka Mineral Resource increased by 0.1 million tonnes owing to additions resulting from drilling and updated metal prices offset by mining depletion and higher NSR COVs. The total Golema Reka Mineral Resource reduced by 0.6 million tonnes mainly owing to a higher NSR COV.

Table 6: Comparison of current and previous Mineral Resource Estimates

Notes

1.        The Svinja Reka Mineral Resources were estimated using the following NSR COVs: 

a.        US$46 per tonne for SLC in 2024; and US$53 per tonne in 2025.

b.        US$53 per tonne for C&F in 2024; and US$65 per tonne in 2025.

c.        US$53 per tonne for LHS in 2024; and US$60 per tonne in 2025.

2.        The Golema Reka Mineral Resources were estimated using a NSR COV of US$53 per tonne in 2024 and US$60 per tonne in 2025.

3.        Metal prices used to estimate the Mineral Resources were:

a.        Zinc - US$2,933 per tonne in 2024; and US$3,041 per tonne in 2025.

b.        Lead - US$2,300 per tonne in 2024; and US$2,506 per tonne in 2025.

c.        Silver - US$26 per ounce in 2024; and US$31 per ounce in 2025.

4.        Contained metal refers to estimated contained metal in the ground not adjusted for metallurgical recovery.

Sasa Ore Reserve Estimate

The audited Ore Reserve statement for Sasa is shown in Table 7.

Table 7: Audited Ore Reserve Statement for Sasa (effective date of 31 December 2025)

Notes

1.        Ore Reserves are reported in accordance with the guidelines of the JORC Code (2012).

2.        Ore Reserves are estimated using NSR cut-off values of US$53 per tonne for SLC, US$65 per tonne for C&F and $60 per tonne for LHS.

3.        Ore Reserves are estimated using metal prices of US$2,644 per tonne for zinc, US$2,179 per tonne for lead and US$27 per ounce for silver.

4.        Ore Reserves are estimated using metallurgical recoveries of 94% for lead, 82% for zinc and 75% for silver.

5.        A minimum mining width of 1.8 metres, including internal but excluding external dilution, has been used.

6.        Contained metal refers to estimated contained metal in the ground not adjusted for metallurgical recovery.

7.        Totals may not represent the sum of the parts owing to rounding.

Material assumptions

Sasa is an operating underground lead-zinc-silver mine focussed on the Svinja Reka deposit. The mine has been in continuous operation since 2006 and costs and revenue are well understood. Modifying factors were based on actual mining, mineral processing, geotechnical, hydrogeological, mining methods, infrastructure, environmental and social, market and economic model information.

A financial model for the life of mine (LoM) showed the Ore Reserves to be economic based on the assumptions used and metal prices of US$2,644 per tonne for zinc, US$2,179 per tonne for lead, and US$27 per ounce for silver.

Ore Reserves are reported for the Svinja Reka deposit only. No Ore Reserves are reported for the Golema Reka deposit because no prefeasibility study or feasibility study has been undertaken.

Classification

The Ore Reserve Estimate is based on the Mineral Resource model for Svinja Reka. Indicated Mineral Resources were converted to Probable Ore Reserves by applying modifying factors. No Proved Ore Reserves were classified because no Measured Mineral Resources were estimated.

Inferred Mineral Resources contained within the stope designs were considered as having zero grade during the stope design process. Inferred Mineral Resources were excluded from the Ore Reserve Estimate, and no value was assigned to them in the economic analysis.

Mining methods and other mining assumptions

Svinja Reka is an operating underground mine and currently produces approximately 800,000 tonnes per annum of ore, with average grades of 3.4% Pb and 2.6% Zn. Mining methods have been developed and upgraded over time as understanding of the orebody geometry and geotechnical properties has increased.

The main access to the underground workings is through the Central Decline, which extends a distance of 3,735 metres from a portal located adjacent to the process plant to a depth of 750 metres Relative Level. Secondary access is provided by the Golema Reka shaft and the Level 14b portal. As of H1 2026, approximately 75% of the ore is hauled by truck and 25% via shaft, with shaft haulage scheduled to be phased out in 2026.

Historically, the mine operated using SLC with 7.0 metre sublevel spacing. As the orebody narrows with depth, the suitability of this method has declined. Since 2023, a transition to LHS with backfill has been ongoing as the primary mining method. C&F is also used in areas requiring greater selectivity or with difficult ground conditions. Both historical voids and newly mined stopes are backfilled using cemented tailings.

Sasa uses conventional trackless underground mining equipment. Ventilation is based on a mechanical overpressure system that supplies fresh air to active mining areas and exhausts return air through a network of raises and airways. Two principal fans are installed, and fresh air is supplied primarily via the decline portals.

Mining dilution is based on mining method, development profile and stope dimensions. Mining recovery is based on operational experience. The dilution and mining recoveries used are 30% dilution and 82% mining recovery for SLC; 10% dilution and 98% mining recovery for C&F; and 18% dilution and 92% mining recovery for LHS. A minimum mining width of 1.8 metres is used.

The current LoM plan extends over approximately nine years. Production is expected to increase to around 830,000 tonnes per annum over the next three years, before reducing in the final year. Average planned grades over this period are 3.5% Pb and 2.5% Zn. Mining is split between LHS (67%), C&F (17%) and SLC (15%).

Processing method and other processing assumptions

The Sasa processing plant was commissioned in 2006 and has a nameplate capacity of 850,000 tonnes per annum of ore. To match the mine production rate, in recent years the plant has operated at throughputs of approximately 800,000 tonnes per annum. The process plant comprises a three-stage crushing circuit and conventional grinding, classification and selective flotation circuits to produce lead and zinc concentrates that are thickened and filtered for shipment by truck.

Metallurgical recoveries used in the LoM plan from 2026 to 2034 are 94% for lead, 86% for zinc and 75% for silver and are consistent with actual recoveries achieved by the plant.

Concentrate grades are approximately 71% Pb in the lead concentrate and 50% Zn in the zinc concentrate. Silver is recovered to both the lead and zinc concentrates but is payable only in the lead concentrate. Silver recovery to the lead concentrate is 75% to a grade of 254 g/t Ag.

No deleterious elements affect the lead concentrate. For the zinc concentrate, no significant deleterious elements result, although it is important to control the silica and iron content, with maximum levels of 2.5% SiO₂ and 11% Fe, respectively.

A paste backfill plant and dry stack plant were constructed and commissioned in 2023 and 2025, respectively. For 2026, it is planned that circa 31% of the tailings will be dry stacked, 36% will be used for paste backfill and 33% stored in the existing TSF4. Paste backfill is the priority method for tailings disposal.

Cut-off grade

Metal prices used in the calculation of NSR for Ore Reserves were US$2,644 per tonne for zinc, US$2,179 per tonne for lead, and US$27 per ounce for silver. Metallurgical recoveries used in the calculation of NSR were 94% for lead, 82% for zinc and 75% for silver. The zinc recovery is conservative and less than the 86% achieved by the processing plant.

The NSR calculation also considered the planned underground mining methods based on orebody geometry and mining level. Between levels 810 metres and 962 metres, and above 990 metres, SLC was used. Between levels 750 metres and 760 metres, and between 800 metres and 810 metres, C&F was used. Below level 750 metres, between levels 760 metres and 800 metres, and between levels 962 metres and 990 metres, LHS was used.

NSR COVs of US$53 per tonne for SLC, US$65 per tonne for C&F and US$60 per tonne for LHS were used to estimate the Ore Reserves.

Estimation methodology

Mine design and scheduling works are carried out using the Deswik Software suite and follow a defined process which conforms to industry best practice for estimating Ore Reserves.

Stope locations are identified using stope optimisations, with final stope designs augmented by manual checks and verifications. Development headings are designed manually according to mining method and location.

Stopes are designed based on average block NSR values which are incorporated into the Mineral Resource block model prior to commencement of the mine design.

Stopes are designed using a combination of optimisation and manual stope/development design. The design basis is driven by the NSR COVs, in conjunction with the geotechnical parameters defining maximum stope dimensions, selected mining method and block value.

Stope optimisation creates a stope shape that maximises the recovered Mineral Resource value above the COV while also allowing for practical mining parameters, including minimum and maximum mining width, anticipated wall dilutions, minimum and maximum wall angles, minimum separation distances between parallel and/or sub-parallel stopes, and minimum and maximum stope heights and widths. Mineable shapes are evaluated using the Mineral Resource block model, based on the NSR COVs and the Mineral Resource classification.

Material modifying factors

An Exploitation Concession (24-5550/1) was granted to Rudnik SASA DOOEL on 13 November 2014, and is valid until 28 September 2030. The Exploitation Concession was expanded in February 2026 from an area of 4.22 square kilometres to approximately 6.87 square kilometres. Rudnik SASA DOOEL can apply to extend the concession beyond the current expiry date. Rudnik SASA DOOEL is a 100% owned subsidiary of CAML.

Permits held by Rudnik SASA DOOEL for the Property are sufficient to ensure that mining activities are conducted within the regulatory framework.

All major infrastructure required by the operation is in place and no additional significant infrastructure projects are currently planned.

The stated Ore Reserves are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues. There is no known mining, metallurgical, infrastructure or other factor that materially affects this Ore Reserve Estimate at this time.

Comparison with previous Ore Reserve Estimate

A previous Ore Reserve Estimate for Sasa was prepared by CAML and reported in accordance with the guidelines of the JORC Code (2012) and with an effective date of 31 December 2024. A comparison of the current and the previous Ore Reserve Estimates is shown in in Table 8.

The total Svinja Reka Ore Reserves reduced by 2.3 million tonnes owing to revisions to the mine design, the application of higher NSR COVs in response to increased assumptions for operating costs; revised assumptions for metal prices and concentrate treatment charges (TCs) and mining depletion.

Table 8: Comparison of current and previous Ore Reserve Estimates

Notes

1.        Ore Reserves were estimated using the following NSR COVs:   

a.        US$46 per tonne for SLC in 2024; and US$53 per tonne in 2025.

b.        US$53 per tonne for C&F in 2024; and US$65 per tonne in 2025.

c.        US$53 per tonne for LHS in 2024; and US$60 per tonne in 2025.

2.        Metal prices used to estimate the Ore Reserves were:

a.        Zinc - US$2,750 per tonne in 2024; and US$2,644 per tonne in 2025.

b.        Lead - US$2,081 per tonne in 2024; and US$2,179 per tonne in 2025.

c.        Silver - US$24 per ounce in 2024; and US$27 per ounce in 2025.

3.        Contained metal refers to estimated contained metal in the ground not adjusted for metallurgical recovery.

Production target

Sasa's life-of-mine plan envisages mining up to approximately 830,000 tonnes annually for eight years commencing 2026.

Material assumptions

The material assumptions underpinning this production target are set out in this section.

The production target is based on Probable Ore Reserves and Inferred Mineral Resources. The Probable Ore Reserves were derived from Indicated Mineral Resources through the application of technical and economic modifying factors. Sasa is operational, and modifying factors were based on operating actuals, as described previously in this announcement. Inferred Mineral Resources were not converted to Ore Reserves. The financial viability of the operation is not dependent on the inclusion of Inferred Mineral Resources in the production schedule.

The Probable Ore Reserves and Inferred Mineral Resources underpinning the production target referred to in this announcement have been prepared by Competent Persons in accordance with the requirements of the JORC Code (2012).

Reserves and Resources underpinning the production target

The relevant proportions of Ore Reserves and Mineral Resources underpinning the production target are as follows:

Table 9: Sasa life-of-mine schedule

Notes

1.        Ore tonnes are Probable Ore Reserves as presented in Section 14 of 'Mineral Resources and Ore Reserves Report, Sasa Mine, North Macedonia' with an effective date of 31 December 2025.

2.        Mining losses and mining dilution applied as detailed in Section 14 of 'Mineral Resources and Ore Reserves Report, Sasa Mine, North Macedonia' with an effective date of 31 December 2025.

3.        Totals may not represent the sum of the parts owing to rounding.

The estimated Ore Reserves and Mineral Resources underpinning the production target have been prepared by Competent Persons in accordance with the requirements of the JORC Code.

Cautionary statement

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 indicated Mineral Resources or that the production target itself will be realised.

Competent Persons

The Mineral Resource Estimate for Kounrad is based on, and fairly represents, information and supporting documentation prepared by Ruslan Erzhanov, an independent Competent Person as defined by JORC. Mr Erzhanov (MSc, FGS, CGeol, PONEN RoK) is a member of the Geological Society of London (UK) and is an employee of SLR Consulting (UK) Ltd. This announcement and the information herein have been reviewed and issued with the prior written consent of Mr Erzhanov as to the form and context in which the Mineral Resource Estimate for Kounrad and the supporting information are presented.

The Ore Reserve Estimate for Kounrad is based on, and fairly represents, information and supporting documentation prepared by Philip King, an independent Competent Person as defined by JORC. Mr King (BSc, ARSM, CEng, FIMMM) is a member of the Institute of Mining, Metallurgy and Materials and is an independent consultant. This announcement and the information herein have been reviewed and issued with the prior written consent of Mr King as to the form and context in which the Ore Reserve Estimate for Kounrad and the supporting information are presented.

The Mineral Resource Estimate for Sasa is based on, and fairly represents, information and supporting documentation prepared by Jason Ché Osmond, an independent Competent Person as defined by JORC. Mr Osmond (BSc, MSc, MCSM, FGS, CGeol, EurGeol) is a member of the Geological Society of London (UK) and is an employee of SLR Consulting (UK) Ltd. This announcement and the information herein have been reviewed and issued with the prior written consent of Mr Osmond as to the form and context in which the Mineral Resource Estimate for Sasa and the supporting information are presented.

The Ore Reserve Estimate for Sasa is based on, and fairly represents, information and supporting documentation prepared by Colin Davies, an independent Competent Person as defined by JORC. Mr Davies (BEng, MSc, CEng, ACSM, MIMMM, QMR) is a member of the Institute of Mining, Metallurgy and Materials and is an employee of SLR Consulting (UK) Ltd. This announcement and the information herein have been reviewed and issued with the prior written consent of Mr Davies as to the form and context in which the Ore Reserve Estimate for Sasa and the supporting information are presented.

This announcement contains inside information for the purposes of Article 7 of Regulation 596/2014. The person responsible for making this announcement is Richard Morgan, Investor Relations Manager.

All dollar amounts in this announcement are US dollars unless otherwise stated.

For further information contact:

Note to editors:

Central Asia Metals, an AIM-quoted UK company based in London, owns 100% of the Kounrad SX-EW copper operation in central Kazakhstan and 100% of the Sasa zinc-lead mine in North Macedonia. The Company also owns an 80% interest in CAML Exploration and 100% in CAML XD, two subsidiaries formed to progress early-stage exploration opportunities in Kazakhstan, and a 32.6% interest in Aberdeen Minerals Ltd, a privately-owned UK company focused on the exploration and development of base metals opportunities in northeast Scotland.

For further information, please visit www.centralasiametals.com and follow CAML on X at @CamlMetals and on LinkedIn at Central Asia Metals Plc