NI43-101 PRELIMINARY ECONOMIC ASSESSMENT OF THE CUKARU PEKI UPPER ZONE DEPOSIT, SERBIA, MARCH 2016

Prepared For

Reservoir Minerals Inc.

Report Prepared by

SRK Consulting (UK) Limited UK6782

SRK Consulting

Cukaru Peki PEA – Details

Important Notice This technical report was prepared by SRK Consulting (UK) Ltd. (“SRK”) for Reservoir Minerals Inc. The information, conclusions and estimates contained herein are based on: i) the information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by Reservoir Minerals Inc. subject to the terms and conditions of its contract with SRK and relevant securities legislation. The contract permits Reservoir Minerals Inc. to file this report as a technical report with the TSX Venture Exchange (the “Exchange”) or with Canadian or similar securities regulatory authorities pursuant to National Instrument 43-101, Standards of Disclosure for Mineral Projects. Except for Exchange purposes or the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk. The responsibility for this disclosure remains with Reservoir Minerals Inc.. The user of this document should ensure that this is the most recent technical report for the property as it is not valid if a new technical report has been issued. This document, as a collective work of content and the coordination, arrangement and any enhancement of said content, is protected by copyright vested in SRK. Outside the purposes legislated under Canadian provincial securities laws and stipulated in SRK’s client contract, this document shall not be reproduced in full or in any edited, abridged or otherwise amended form unless expressly agreed in writing by SRK. © 2016 SRK Consulting (UK) Ltd.

SRK Legal Entity:

SRK Consulting (UK) Limited th

SRK Address:

5 Floor Churchill House 17 Churchill Way City and County of Cardiff, CF10 2HH Wales, United Kingdom. March 2016

Date: Project Number:

UK6782

SRK Project Director:

Mark Campodonic

SRK Project Manager:

Martin Pittuck

Principal Consultant (Resource Geology) Corporate Consultant & Director Reservoir Minerals Inc

Client Legal Entity: Client Address:

U6782 Cukaru Peki PEA Report v30.docx

Suite 501 543 Granville Street Vancouver BC Canada V6C 1X8

March 2016

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Cukaru Peki PEA – Details

Report Title:

NI43-101 PRELIMINARY ECONOMIC ASSESSMENT OF THE CUKARU PEKI UPPER ZONE DEPOSIT, SERBIA, MARCH 2016 st

Effective Date:

31 March, 2016

Signature Date

13 April, 2016

Project Number: Qualified Persons:

th

UK6782 Martin Pittuck, Corporate Consultant (Mining Geology) Tim McGurk, Corporate Consultant (Mining Engineering)

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SRK Consulting (UK) Limited 5th Floor Churchill House 17 Churchill Way City and County of Cardiff CF10 2HH, Wales United Kingdom E-mail: [email protected] URL: www.srk.co.uk Tel: + 44 (0) 2920 348 150 Fax: + 44 (0) 2920 348 199

NI43-101 PRELIMINARY ECONOMIC ASSESSMENT OF THE CUKARU PEKI UPPER ZONE DEPOSIT, SERBIA, MARCH 2016 1

EXECUTIVE SUMMARY

1.1

Introduction Reservoir Minerals Inc, listed on the Toronto Venture Stock Exchange as RMC, (“RMC”, “Reservoir” or “the Client”) currently owns a 45% interest in the Timok Project comprising four 2 exploration Permits in eastern Serbia, with a total area 212.58 km . The Cukaru Peki coppergold deposit was discovered in 2012 and is located within the Brestovac-Metovnica Exploration Permit, which was the subject of a maiden NI 43.101 Mineral Resource estimate prepared by SRK Consulting (UK) Ltd (“SRK”) in January 2014. During 2014 and 2015, further drilling and various early stage technical and environmental studies have been undertaken which are reviewed and incorporated in this Mineral Resource update and Preliminary Economic Assessment (“PEA”) of the Cukaru Peki deposit (the “Project”). The PEA is preliminary in nature and it includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as Mineral Reserves. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. There is no certainty that the PEA will be realised. The Project area is located in eastern Serbia on a gently rolling plateau between 300 and 400 metres above mean sea level (“amsl”) and has a moderate-continental climate. It is located some 5 km south of the town of Bor, which is a regional administrative and mining centre, located approximately 245 km by road southeast of Belgrade, the capital of Serbia. The site is favourably located for mining infrastructure (road, rail, power, water) and near by the recently upgraded copper smelter complex in Bor. The operator of the Project is Freeport-McMoRan Exploration Corporation (“Freeport”, or “FMEC”). After acquiring 55% equity interest under the Rakita Agreement, Freeport gave notice to Reservoir in July 2012 that it had elected to sole fund expenditures on or for the benefit of the Project until the completion and delivery to the Company of a feasibility study to bankable standards (the “Bankable Feasibility Study”), subject to its right to cease such funding at any time. The Bankable Feasibility Study must be in such form as is normally required by substantial, internationally recognized financial institutions for the purpose of deciding whether or not to loan funds for the development of mineral deposits. If Freeport completes the Bankable Feasibility Study, Freeport will indirectly own 75% and Reservoir 25% of the Project.

Registered Address: 21 Gold Tops, City and County of Newport, NP20 4PG, Wales, United Kingdom. SRK Consulting (UK) Limited Reg No 01575403 (England and Wales) Page 1 of 268

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Cukaru Peki PEA – Executive Summary

Reservoir and Freeport hold their respective indirect interest in the Project via a wholly owned local Serbian registered company, Rakita Exploration d.o.o (“Rakita”), which owns the Project exploration permits comprising the Jasikovo-Durlan Potok, Brestovac-Metovnica, Leskovo and the recently awarded Brestovac Zapad (“Brestovac West”) Exploration Permits which are all valid until February 2017, with the exception of the Brestovac Zapad Exploration permits which is valid until April 2018. The total area of the exploration permits is 212.58 square kilometres. Exploration permits are renewable subject to completion of agreed work programs and approval of the Serbian Ministry of Mining and Energy. On 3 March 2016, Global Reservoir Minerals (“BVI”) Inc, a wholly owned subsidiary of Reservoir, received a notice of sale and offer from Freeport International Holdings (BVI) Ltd. (“Freeport”) (the “Notice of Sale and Offer”), wherein Freeport: (i) provided notice to Reservoir of the proposed sale to Lundin Mining Corporation (“Lundin”) of an interest in Freeport International Holdings (BVI) Ltd, the entity through which Freeport holds its interest in the Project, under a Joint Venture/Shareholders Agreement dated December 15, 2015 among Freeport, Reservoir and Timok JVSA (BVI) Ltd. (the “Joint Venture Agreement”); and (ii) offers to sell to Reservoir on the same terms and conditions as those agreed with Lundin pursuant to Reservoir’s right of first refusal under Section 15.04 of the Joint Venture Agreement. Reservoir has until 3 May 2016 to decide whether it will exercise its right of first refusal.

1.2

Geology and Resource Geologically, the region is part of the Western or Eurasian Tethyan Belt, which hosts Mesozoic to Cenozoic subduction-related gold and base metal deposits from Romania and Serbia through to Turkey, Iran and Pakistan. Locally, the Timok Metallogenic Complex has one of the highest concentrations of copper enrichment in the Belt, containing world-class examples of porphyry and high sulphidation epithermal types of copper-gold deposits, such as Bor and Majdanpek, which have an estimated historical production of some six million tonnes of copper and nearly ten million ounces of gold. The Bor cluster of deposits is hosted by Upper Cretaceous andesites and volcaniclastics that continue at least five kilometres south to Cukaru Peki, where the Cretaceous is overlain by a Miocene basin containing clastic sediments that can reach up to 600 m in thickness. The Cukaru Peki deposit is subdivided into an Upper Zone (“UZ”) of high-sulphidation epithermal mineralization and the underlying Lower Zone (“LZ”) of porphyry type mineralization which has not yet been modelled due to insufficient drill data and geometrical understanding, and is not included in current the resource estimates. Despite incomplete drill coverage in the Lower Zone, there are indications of a large deposit that may be amenable to a bulk mining method. The Lower Zone is currently an Exploration Target that remains open to the north and at depth below the UZ, with several drillhole intersections that outline the potential magnitude of the mineralised system.

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Cukaru Peki PEA – Executive Summary

The close spatial association of the high sulphidation and porphyry copper-gold mineralization at Cukaru Peki is very similar to that observed and described from the active Bor mining district, and Company geologists interpret the LZ and UZ mineralization at Cukaru Peki to be comparable to, respectively, the porphyry and high sulphidation mineralization in the Bor District. The porphyry style mineralization at Bor includes the Borska Reka deposit (see the RTB Bor website, www.rtb.rs, and Petrović et al., 2012) and the spatially associated high sulphidation copper gold mineralization at Bor which occurs in several deposits (Tilva Ros, Tilva Mika, Coka Dulkan, etc.) that have been mined out with past production estimated as approximately 200 Mt at 1.5% copper and 0.8 g/t gold; (see BRGM RC-51448-FR,). The Company considers that the mineralisation styles in the Bor District are relevant to the assessment of the Timok Project; however, it should be noted that mineralisation mined in other deposits in the Bor District may not necessarily be indicative of mineralisation at Cukaru Peki. Drilling completed on the Cukaru Peki Upper Zone (“UZ”) deposit between 2012 and 2016 has intersected high sulphidation epithermal (“HSE”) pyrite and copper sulphide-rich mineralisation at depths of between 400 m and 1,000 m below surface. The highest grades are at the top of the body where a lens of high grade massive pyrite breccia has been outlined which is expected to be sold directly to smelters without any upgrading required. Beneath this zone, a semi-massive sulphide zone extends downwards within altered andesite host rock; copper and gold grades decrease downwards. At the Cukaru Peki Upper Zone deposit, 111 drillholes totalling some 90,739 m have been drilled. Excluding 14 non-sampled hydrogeological holes, most of these have been sampled and assayed using industry standard methods, except 22 drillholes (17,074 m) which have not yet been assayed; visual copper grade estimates in these have been used to guide geological modelling, but these have not been used for grade estimation in the block model that has been used to derive the Mineral Resource and has been used for the mining study in the Preliminary Economic Assessment (“PEA”). Drillholes have variable spacing and orientation (providing intersections spaced between 25-100 m apart) which results in the majority of the Mineral Resource being classified as “Inferred”; it is expected that further drilling towards a 25 m grid coverage will achieve the confidence necessary to convert more of the deposit to Indicated and possibly Measured Mineral Resources which would then allow Mineral Reserves to be derived. On a 100% attributable basis, the Cukaru Peki Upper Zone Mineral Resource comprises an Indicated Mineral Resource of some 1.7 Mt above a cut-off grade of 0.75% copper, with average grades of 13.5% copper, 10.4 g/t gold and 0.23% arsenic and an Inferred Mineral Resource of some 35.0 Mt above a cut-off grade of 0.75% copper, with average grades of 2.9% copper, 1.7 g/t gold and 0.17% arsenic. The Table ES1-1 gives the current Mineral Resource Estimate which has been prepared in accordance with the CIM Code.

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Table ES1-1: Category Indicated Inferred

Cukaru Peki PEA – Executive Summary

SRK Mineral Resource Statement as at 31 March 2016 for the Cukaru Peki Deposit Upper Zone Grade Category %Cu

Tonnes Mt

Grade

Metal % Cu

g/t Au

% As

Cu Mt

Au Moz

>10.0

1.2

15.5

12.4

0.21

0.2

0.5

0.75-10.0

0.5

8.2

5.1

0.27

0.04

0.08

>10.0 0.75-10.0

Total-Indicated Total-Inferred

1.2

13.2

10.5

0.20

0.2

0.4

33.8

2.5

1.4

0.17

0.8

1.5

1.7 35.0

13.5 2.9

10.4 1.7

0.23 0.17

0.2 1.0

0.6 1.9

1. The cut-off grade used for the estimate is 0.75% Cu. 2. All figures are rounded to reflect the relative accuracy of the estimate. 3. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. 4. The Mineral Resource is reported on 100% basis; currently 45% is attributable to Reservoir.

The majority of the estimate is an Inferred Mineral Resource, which SRK expects can be upgraded to an Indicated Mineral Resource with further drilling and sampling. There are no Mineral Reserves at this time.

1.3

Mining The deposit is likely to be mined by conventional underground mining methods in two or three phases: Phase 1 “DSO” Starter Mine will access very high grade mineralisation (13% Cu and 10.6 g/t Au) which may be shipped directly to smelters after crushing and grinding only. The DSO mineralisation is at the top of the deposit, some 400 m below surface and will be mined at a rate of up to 0.6 Mtpa for 3.5 years; initial mining access is via a 2.5 km long dual decline. Phase 2 Upper Zone (UZ) Main Mine is envisaged to extract the remaining deposit down to a depth of some 800 m below surface where copper grades reduce to 2.5%. Production increases over a 10 year period from 0.8 to 2.0 Mtpa. The PEA base case LOM is based on Phases 1 and 2 only. Phase 3 Upper Zone Main Mine Extension which would extract the remaining deposit until copper grades fall to 1.0%. Production continues for another 8 years at a rate of 2.0 Mtpa All phases of mining are planned to use Sub Level Open Stoping (“SLOS”) mining method with back fill. This approach reduces initial development capital and maximises early copper production and revenue. This is a key assumption to achieve the high economic returns of the Project. Mining production grade and tonnage profiles have been generated using for each mine phase (DSO and UZ) by scheduling production from basic stope designs based on Indicated and Inferred Mineral Resources identified in SRK’s block model using industry standard software and methods. The annual production target used for DSO mine planning was 70 kt contained copper in the mill feed. Using a diluted cut-off grade of 9.3% Cu, it was determined that approximately 1.9 Mt of material at an average mill feed grade of 13.0% copper and 10.6 g/t gold will acheive this objective, containing a total of 248 kt of copper and 649 koz of gold. This equates to approximately 545,000 tpa of DSO, for 3.5 years. This material is sourced from both Indicated and Inferred Mineral Resources.

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Cukaru Peki PEA – Executive Summary

To ensure continuous production from the underlying UZ main mine, a flotation plant is scheduled for commissioning 3.5 years after DSO production commences. The float material available at a 1.0% Cu cut-off grade is 27.5 Mt at an average grade of 3.0% copper and 1.75 g/t gold. In the first six years, mill feed grades are much higher which allows the plant to process 0.8 Mtpa recovering >70 ktpa copper and 150 koz gold. After this period, mill feed grades are lower so the plant capacity is increased to 2.0 Mtpa and the annual contained copper reduces gradually to 20 ktpa and gold production reduces gradually to 20 kozpa reflecting the declining copper and gold grades with depth.

1.4

Processing The PEA assumptions for concentrate production are based on preliminary metallurgical testwork and mineralogical work completed on Cukaru Peki UZ samples. The copper mineralogy consists primarily of covellite with lesser enargite, and minor colusite, bornite and chalcopyrite. Enargite and colusite contain arsenic and therefore the arsenic grade of any copper concentrate will be a key consideration for smelter choice, terms and penalty charges. The PEA envisages the following production schedule: •

DSO with copper grade of 12-14%, a gold grade of 9-15 g/t and an arsenic grade of 0.13 – 0.25% that could be smelted directly after crushing and milling during the first three to four years of production;

•

Subsequent production of two copper concentrates: o

Up to 100 ktpa of a ‘Low Arsenic’ high Cu concentrate with a copper grade of 5258%, a gold grade of 3.9 – 6.3 g/t and an arsenic grade of 0.48% after blending; and

o

18 – 52 ktpa of a ‘High Arsenic’ relatively high Cu concentrate with a copper grade of 37%, a gold grade of 6.2 – 12.8 g/t and an arsenic grade of 2.5 – 5.5% after blending.

Both concentrates are within marketable ranges, with no other significant deleterious elements identified. Table ES1-2 summarises value drivers through the mine phases. Phase 1 DSO production contains approximately a third of the gold and copper in the Base Case life of mine mill feed (“ROM”), which will derive revenue in the first four years of the mine life according to the smelter payment terms assumed in the PEA. Once concentrate production commences 60% of the copper in the Base Case life of mine production is recovered in concentrate. Just over half of the gold in ROM reports to a pyrite concentrate which is considered as a waste product in the PEA; this will be stored separately and will be targeted for additional processing testwork in future technical studies. Table ES1-2:

Mine Gate Production (Base Case) Metal in ROM

Mining Period

Year 1 - 19

Recovered Metal sent to smelter "ex mine gate" Combined low and high DSO arsenic concentrates Year 1 - 4 Year 4 - 12

Copper (kt)

858

248

507

Gold (koz)

1,875

649

184

Arsenic (kt)

33

4

18

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On-site operating cost assumptions are based on relevant industry benchmarks. Capital costs for the concentrator are also benchmarked against industry peers, scaled according to production rate.

1.5

Infrastructure Surface infrastructure requirements that have been considered in this study include power and water supply, access, administration facilities and export logistics. The Project site is close to existing road, rail and power, there is skilled labour available with mining experience from the regional capital Bor which is 5 km from the UZ Mine. The government owned RTB Bor smelter, also in Bor town, is anticipated to have capacity to smelt and refine some of the mine’s product. Supply of water for the mine is not expected to be a concern with opportunities being considered for nearby rivers and Bor Lake. Several tailings disposal sites have been assessed and costs of impoundment construction have been estimated to a basic level. It is assumed that the tailings impoundment construction material and mining backfill aggregate will be available from a supplier or from a quarry close to the mine though this has not been studied in detail for this study. Consideration has been given to a number of possible options for transporting product locally to the Bor smelter complex and export to European and other international smelters making use of local rail head options to a Danube port or a coastal port to ensure export options are reasonably well understood.

1.6

Marketing Cliveden Trading AG (“CTAG”) contributed a marketing study covering post-mine-gate costs (freight from mine gate to smelters, treatment, penalty element and refining charges) using current data from a range of smelters capable of taking DSO and concentrate products. Particular attention was given to a number of possible deleterious elements; at this stage, only arsenic grades are expected to impact the net revenues and the overall sales strategy. Consideration was given to Bor itself and other smelters. Cost parameters for the economic analysis are based on anticipated smelter capacity and terms in relevant geographical locations based on CTAG’s analysis of historical and current market conditions with an assessment of planned / anticipated developments that are likely in the coming years.

1.7

Environmental Considerations The Project site is in the same municipality at the RTB Bor operations (in the Bor Muncipality) and is about 8 km south of the RTB Bor operations. The towns of Brestovac and Slatina and a small airport lie close to possible Project infrastructure sites. Project infrastructure would largely be located on land used for agriculture. The Project is in the Timok River basin. The Timok River is a transboundary river which flows along the border of Serbia and Bulgaria for 17.5 km before its confluence with the Danube River and has been polluted by the RTB Bor copper mining and smelting operations since they began in 1900. These operations are state owned and the Serbian government has initiated a regional project that includes rehabilitation of the areas degraded by the mining operations. Air pollution by the RTB Bor smelting facilities has also been an issue in the region for decades but is now being addressed by replacement of these facilities with a new smelter and acid production plant designed to meet European emission and other standards.

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Cukaru Peki PEA – Executive Summary

The RTB Bor environmental and social legacy will have consequences for the Project. Regulatory authorities and other stakeholders are likely to subject the Project to scrutiny during permitting processes. To get approvals for the Project, it will be necessary to ensure that environmental and social impacts are properly defined and appropriate measures are put in place to avoid these or, at least, to reduce these to acceptable levels. Constructive relationships with stakeholders, established by means of effective stakeholder engagement during the planning of the project, will facilitate approval of the Project. To date, limited stakeholder engagement and environmental and social baseline studies have been undertaken for the Project by FMEC. Environmental approvals required to develop the mine include: •

the primary environmental approval, which is EIA approval;

•

approval from the ministry responsible for water resources;

•

approvals from planning authorities; and

•

approvals from the Institute for Nature Conservation and the Institute for Protection of Cultural Heritage.

Reservoir has allowed reasonable timeframes in the Project development schedule to obtain the environmental approvals for the development of the mine. There is a need to develop a better understanding of the environmental and planning approvals required for the bulk sampling works and building of the DSO plant. To develop this understanding, it will be necessary to consult with relevant regulatory authorities and share details on the nature and scale of the proposed developments during these consultations. Reservoir has allowed a period of 12 months to apply for and obtain the necessary approvals. Surface rights, in the form of land ownership or easement agreements, will be required to develop the project surface infrastructure and to complete mine permitting. The land in the project area is divided into hundreds of small portions of land which will have to be acquired to develop the mine. The land above the deposit on which the airport is situated will likely need to be purchased through municipal auction. A land acquisition plan for the Project has been developed and initiated using local and international consultants. At the end of 2015, 71 parcels covering 131 ha had been purchased or were in advanced negotiation; these include significant buildings such as the current project office, core shed, storage yard and an ex-military barracks. The surrounding land is mostly agricultural, so no villages will need to be moved and the number of inhabited dwellings in the area is small. The plan will need to address physical and economic displacement that could arise from the land acquisition. The updated mining law now includes rights for the government to acquire land via court order if necessary.

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1.8

Cukaru Peki PEA – Executive Summary

Capital and Operating Costs Costs have been estimated to a basic level of accuracy. They include an extra 30% indirect cost on construction items to cover engineering, procurement and construction management and a subsequent 25% contingency amount to account for the preliminary nature of these cost estimates. Capital and operating costs have been derived from a variety of methods mainly based on benchmarking using SNL Raw Materials (“SNL”) and Thomson Reuters GFMS (“TR”) databases with some costs built up from basic first principles where possible. A key consideration in the PEA is the establishment of a relatively low cost and high margin DSO Starter Mine which can be built for an estimated capital investment of USD213M including the initial decline access. DSO production is expected to generate sufficient profit to fund the additional capital required to ramp up to the eventual 2 Mtpa UZ Main Mine production rate in two stages over six years which, in addition to sustaining capital costs amounts to USD226M for the base case mine. A closure cost of USD23M has been provisionally estimated. Operating costs used in the PEA for the initial DSO Mine are USD63/t reflecting the low production rate; these gradually reduce to USD40/t for the steady state 2 Mtpa UZ Mine. All operating costs compare well with industry benchmarks.

1.9

Development Cost and Schedule The development schedule and costs presented in this PEA are preliminary. The 3-year timeframe required to secure the necessary permits is considered reasonable assuming the owners receive support and positive engagement with the relevant authorities and successful execution of significant exploration programmes. Completion of technical studies including all aspects of permitting will need to be carefully managed in order to proceed at a rate and cost assumed in the PEA. Land acquisition for the UZ is expected to cost some USD10 – 20M. This includes finalising the current south portal acquisition process and allowances for the north portal option, the airport, UZ mill site and UZ tailings facilities. Additional land acquisition for the LZ footprint including expanded tailings storage and disturbance area would likely cost a further USD10 - 20M. In addition to land acquisition, the necessary exploration and development drilling programmes and associated technical and environmental studies are expected to cost some USD50 – 70M. This includes allowances for infill drilling of the UZ and 200 x 200 m drilling of the LZ, geotechnical drilling for the mine and decline access route and USD25M for technical studies. Land acquisition and technical programmes will need to be carefully managed to feed into the permitting of the exploration decline by May 2017, and into the finalisation of Mine Use Permitting by Q1 2019 in order to commence production in mid-2019 as envisaged in the PEA.

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Cukaru Peki PEA – Executive Summary

1.10 Economic Analysis A technical-economic model has been developed to provide a preliminary economic analysis of the potential viability of mineral resources. This considers operating costs (mining, processing, general and administration, concentrate freight and insurance costs, smelter treatment and refining charges), capital costs (mine access, mining equipment, mine infrastructure, backfill plant, process plant, surface infrastructure, tailings dam, indirect costs and contingency) and revenues (based on a copper price of USD3/lb, a gold price of USD1,200/oz, mineral processing recoveries to concentrate, smelter payability terms, arsenic penalties, royalty deductions and corporate tax). Copper accounts for 70% of revenue during the initial DSO mining phase, and 95% in subsequent mining phases when concentrates are produced. Gold is a significant contributor to early DSO revenue, but less significant for revenue derived from concentrates. Copper mined and fed to the mill is in the range of 65 to 85 ktpa. Net smelter return per tonne of copper fed to the mill (after credit for gold) starts in year 1 at USD 6,770/t Cu and progressively reduces to a low of USD4,000/t Cu at the end of the mine life. Steady state operating cost per milled tonne is USD 80 to 90/t for production of DSO reducing to USD 40 to 50/t in later years. Three Phases were modelled, in summary: •

The DSO Starter Mine, in isolation, has a post-tax NPV of USD 679 M with an IRR of 103%.

•

Additional concentrate production from the base case UZ Main Mine (stopping when grade falls to 2.5% CU) increases the post-tax net present value (“NPV”) to USD1,552M and the internal rate of return (“IRR”) is 106%.

•

Additional concentrate production from UZ Main Mine Extension (stopping when grades fall to 1.0% copper), increases the post-tax NPV to USD1,631M and the IRR remains at 106%.

1.11 Recommendations Overall, the Project has very strong economic returns and SRK recommends that the Project proceeds towards pre-feasibility study. The potential to establish production in three years is realistic, as long as a number of technical and environmental studies are commissioned in Q2 2016 and as long as interactions with local and governmental stakeholders and regulatory authorities are proactively and sensitively managed.

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1.12 Abbreviations Abbreviation amsl

Meaning

APB

andesite pyrite Bbeccia

ARD

acid rock drainage

ARDML

acid rock drainage metal leaching

AWD

accelerated weigth drop

CAF

cemented agreggate fill

CBMB

carpathian-balkan metallogenic belt

CTAG

Cliveden Trading AG

HSE

high sulphidation epithermal

HSF

hydraulic sand fill

IDW

inverse distance weighting

IP

induced polarisation

IRR

internal rate of return

LZ

lower zone

NPV

net present value

NSR

net smelter return

OK

ordinary kriging

PDEX

Phelps Dodge Exploration Corp

PEA

Preliminary Economic Assessment

QKNA

quantitative kriging neighbourhood analysis

QP

Qualified Person

ROM

run of mine

SEE

Southeast Europe Exploration d.o.o

SLOS

sub-level open stoping

TMC

Timok metallogenic complex

UZ

upper zone

above mean sea level

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Table of Contents 1

EXECUTIVE SUMMARY ..................................................................................... 1 1.1 Introduction .............................................................................................................................. 1 1.2 Geology and Resource ............................................................................................................ 2 1.3 Mining ...................................................................................................................................... 4 1.4 Processing ............................................................................................................................... 5 1.5 Infrastructure ............................................................................................................................ 6 1.6 Marketing ................................................................................................................................. 6 1.7 Environmental Considerations ................................................................................................. 6 1.8 Capital and Operating Costs ................................................................................................... 8 1.9 Development Cost and Schedule ............................................................................................ 8 1.10 Economic Analysis ................................................................................................................... 9 1.11 Recommendations ................................................................................................................... 9 1.12 Abbreviations ......................................................................................................................... 10

2

INTRODUCTION ............................................................................................... 23 2.1 Background ............................................................................................................................ 23 2.2 Terms of Reference ............................................................................................................... 23 2.3 Work Completed .................................................................................................................... 24 2.4 Scope of Work ....................................................................................................................... 24 2.5 Sources of Information........................................................................................................... 25 2.6 Requirement, Structure and Compliance .............................................................................. 25 2.7 Details of Personal Inspections ............................................................................................. 25 2.8 Limitations, Reliance on SRK, Declaration, Consent, Copyright ........................................... 26

3

RELIANCE ON OTHER EXPERTS ................................................................... 27

4

PROPERTY DESCRIPTION AND LOCATION ................................................. 28 4.1 Location ................................................................................................................................. 28 4.2 Property Description .............................................................................................................. 30 4.2.1 Mineral Tenure in Serbia ............................................................................................. 31 4.2.2 Exploration Permit Ownership ..................................................................................... 32 4.2.3 Exploration Permit Title ............................................................................................... 32 4.2.4 State Royalties and Royalty Agreements .................................................................... 35 4.3 Permitting Requirements for the Project................................................................................ 37 4.3.1 Mining Permits to be Obtained .................................................................................... 37 4.4 Development Schedule.......................................................................................................... 41 4.5 Other Development Permits .................................................................................................. 43 4.5.1 Development of an Exploration Decline ...................................................................... 43 4.5.2 Construction of DSO Process Plant and associated Surface Infrastructure ............... 43 4.6 Surface Rights & Land Acquisition ........................................................................................ 43 4.7 Project Development Costs ................................................................................................... 44 4.8 Environmental Considerations ............................................................................................... 45

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4.8.1 Closure ........................................................................................................................ 45

5

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY .............................................................................................. 46 5.1 Accessibility ........................................................................................................................... 46 5.2 Climate ................................................................................................................................... 46 5.3 Infrastructure and Local Resources ....................................................................................... 46 5.4 Physiography ......................................................................................................................... 47 5.5 Seismicity ............................................................................................................................... 48

6

HISTORY ........................................................................................................... 49 6.1 Introduction ............................................................................................................................ 49 6.2 History of Exploration and Mining .......................................................................................... 49 6.2.1 Historical Exploration and Mining to 2004 ................................................................... 49 6.2.2 Exploration 2004 – 2010 ............................................................................................. 50 6.3 Historical Estimates ............................................................................................................... 50 6.4 Historical Production .............................................................................................................. 50

7

GEOLOGICAL SETTING AND MINERALISATION .......................................... 51 7.1 Regional Geology .................................................................................................................. 51 7.2 Permit/ Local Geology ........................................................................................................... 52 7.3 Structural Geology ................................................................................................................. 54 7.4 Deposit Geology .................................................................................................................... 55 7.4.1 Introduction .................................................................................................................. 55 7.4.2 Lithology ...................................................................................................................... 55 7.4.3 Alteration ..................................................................................................................... 55 7.4.4 Mineralisation .............................................................................................................. 58 7.4.5 Structural Geology ....................................................................................................... 60

8

DEPOSIT TYPES .............................................................................................. 62 8.1 Mineralisation in the Bor District ............................................................................................ 62 8.2 Other Analogues .................................................................................................................... 63

9

EXPLORATION ................................................................................................. 65 9.1 Historical Exploration ............................................................................................................. 65 9.2 Exploration by Rakita ............................................................................................................. 65

10 DRILLING .......................................................................................................... 67 10.1 Historical Drilling Programmes .............................................................................................. 67 10.2 Current Drilling Programmes ................................................................................................. 67 10.2.1 Introduction .................................................................................................................. 67 10.2.2 Summary of Data Quantity .......................................................................................... 67 10.2.3 Collar Surveys ............................................................................................................. 68 10.2.4 Downhole Surveys....................................................................................................... 69 10.2.5 Hole Orientation........................................................................................................... 69 10.2.6 Diamond Drilling Procedure ........................................................................................ 69 U6782 Cukaru Peki PEA Report v30.docx

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10.2.7 Core Recovery............................................................................................................. 71 10.2.8 Core Storage ............................................................................................................... 71 10.3 SRK Comments ..................................................................................................................... 71

11 SAMPLE PREPARATION, ANALYSIS AND SECURITY ................................. 72 11.1 Introduction ............................................................................................................................ 72 11.2 Diamond Drilling Sample Preparation and Chain of Custody................................................ 72 11.3 Sample Preparation and Analysis ......................................................................................... 72 11.4 Specific Gravity Data ............................................................................................................. 73 11.5 SRK Comments ..................................................................................................................... 74

12 DATA VERIFICATION ....................................................................................... 75 12.1 Verifications by SRK .............................................................................................................. 75 12.1.1 Verification of Sample Database ................................................................................. 75 12.2 Verifications by Reservoir ...................................................................................................... 75 12.3 QAQC for Copper, Gold and Arsenic 2011-2015 .................................................................. 75 12.3.1 Standards .................................................................................................................... 75 12.3.2 Blanks .......................................................................................................................... 76 12.3.3 Duplicates .................................................................................................................... 76 12.4 SRK Comments ..................................................................................................................... 77

13 MINERAL PROCESSING AND METALLURGICAL TESTING ......................... 81 13.1 Introduction ............................................................................................................................ 81 13.1.1 Composite Sample Description ................................................................................... 81 13.1.2 Test program overview ................................................................................................ 83 13.2 FMC Testwork ....................................................................................................................... 84 13.3 Mineralogical Results ............................................................................................................. 85 13.3.1 Qualitative X-Ray Diffraction Analysis ......................................................................... 85 13.3.2 Theoretical Copper Grade and Recovery ................................................................... 86 13.4 Metallurgical Testwork Results .............................................................................................. 86 13.4.1 Rougher Flotation ........................................................................................................ 87 13.4.2 Cleaner Flotation ......................................................................................................... 87 13.4.3 Locked Cycle Tests ..................................................................................................... 88 13.4.4 Cyanidation Bottle Roll ................................................................................................ 92 13.4.5 Grindability Tests ......................................................................................................... 92 13.4.6 DSO and Concentrate Multi Element Assays ............................................................. 93 13.5 Metallurgical Recovery .......................................................................................................... 95 13.6 SRK Comments ..................................................................................................................... 95

14 MINERAL RESOURCE ESTIMATE .................................................................. 97 14.1 Introduction ............................................................................................................................ 97 14.2 Resource Estimation Procedures .......................................................................................... 97 14.3 Resource Database ............................................................................................................... 97 14.4 Statistical Analysis – Raw Data ............................................................................................. 97 U6782 Cukaru Peki PEA Report v30.docx

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14.5 3D Modelling .......................................................................................................................... 98 14.5.1 Geological Wireframes ................................................................................................ 98 14.5.2 Mineralisation Wireframes ........................................................................................... 99 14.5.3 Mineralisation Model Coding ..................................................................................... 102 14.6 Compositing ......................................................................................................................... 103 14.7 Evaluation of Outliers........................................................................................................... 103 14.8 Geostatistical Analysis ......................................................................................................... 104 14.9 Block Model and Grade Estimation ..................................................................................... 105 14.10 Final Estimation Parameters ............................................................................................. 106 14.11 Model Validation and Sensitivity ....................................................................................... 107 14.11.1 Sensitivity Analysis ............................................................................................... 107 14.11.2 Block Model Validation ......................................................................................... 107 14.12 Mineral Resource Classification ........................................................................................ 111 14.13 Mineral Resource Statement............................................................................................. 113 14.14 Cut-off Grade Sensitivity Analysis ..................................................................................... 113 14.15 Comparison to Previous Mineral Resource Estimates...................................................... 114 14.16 Exploration Potential ......................................................................................................... 115

15 MINERAL RESERVE ESTIMATES ................................................................. 116 16 MINING METHODS ......................................................................................... 117 16.1 Geotechnical Considerations ............................................................................................... 117 16.2 Hydrogeology and Mine Dewatering ................................................................................... 120 16.2.1 Groundwater Model ................................................................................................... 121 16.2.2 SRK Comments ......................................................................................................... 122 16.2.3 Recommendations..................................................................................................... 122 16.3 Mining Method Selection ..................................................................................................... 123 16.4 Mine Extraction Design ........................................................................................................ 123 16.4.1 Stope Design ............................................................................................................. 123 16.5 Access Options .................................................................................................................... 129 16.5.1 Production Shaft ........................................................................................................ 129 16.5.2 Conveyor in 1:5 decline ............................................................................................. 129 16.5.3 Truck haulage in 1:7 decline ..................................................................................... 129 16.5.4 Decline Location ........................................................................................................ 131 16.6 Mine Development ............................................................................................................... 132 16.6.1 Decline Development ................................................................................................ 132 16.6.2 Mine Development..................................................................................................... 132 16.6.3 Development Schedule ............................................................................................. 135 16.6.4 Ground Support ......................................................................................................... 137 16.6.5 Dewatering ................................................................................................................ 137 16.6.6 Ventilation .................................................................................................................. 138 16.7 Mine Extraction design and Sequencing ............................................................................. 139

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16.7.1 LoM Production Scenarios ........................................................................................ 139 16.7.2 Dilution and Recovery ............................................................................................... 140 16.7.3 Production Schedule ................................................................................................. 142 16.8 Operating Strategy............................................................................................................... 143 16.8.1 Overall Extraction Strategy ........................................................................................ 143 16.8.2 Haulage Equipment ................................................................................................... 143 16.8.3 Other Mining Equipment ............................................................................................ 145 16.8.4 Stope Drilling and Extraction ..................................................................................... 145 16.8.5 Backfill ....................................................................................................................... 147 16.8.6 Mine Personnel.......................................................................................................... 151 16.8.7 Mining Risks .............................................................................................................. 151 16.8.8 Mining Opportunities.................................................................................................. 152 16.8.9 Mining Recommendations ......................................................................................... 153

17 RECOVERY METHODS .................................................................................. 154 17.1 Introduction .......................................................................................................................... 154 17.2 Cukaru Peki Facility Description .......................................................................................... 155 17.2.1 Capital and Operating Cost estimation...................................................................... 155 17.3 Bor Smelter Facility Description .......................................................................................... 156 17.4 XGC Smelter Facility Description ........................................................................................ 156 17.5 Brief Summary of Processing Alternatives .......................................................................... 157 17.5.1 Option for a Third Concentrate .................................................................................. 157 17.5.2 Hydrometallurgical Plant ........................................................................................... 157

18 PROJECT INFRASTRUCTURE ...................................................................... 159 18.1 Introduction .......................................................................................................................... 159 18.2 Mine Site Surface Infrastructure .......................................................................................... 159 18.2.1 Power Infrastructure .................................................................................................. 159 18.2.2 Potable Water and Sewerage Infrastructure ............................................................. 160 18.2.3 Surface Layout .......................................................................................................... 160 18.3 Local and Regional Export Logistics ................................................................................... 160 18.3.1 Introduction ................................................................................................................ 160 18.3.2 Export Ports ............................................................................................................... 160 18.3.3 Transportation to Ports .............................................................................................. 161 18.3.4 Road/Rail Transfer Station ........................................................................................ 161 18.3.5 Road Transportation to Rail Transfer Station ............................................................ 162

19 MARKET STUDIES AND CONTRACTS ......................................................... 163 19.1 Marketing Studies ................................................................................................................ 163 19.1.1 Introduction ................................................................................................................ 163 19.1.2 Payment Terms ......................................................................................................... 164 19.1.3 DSO Market ............................................................................................................... 164 19.1.4 Concentrate Market ................................................................................................... 165 U6782 Cukaru Peki PEA Report v30.docx

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19.1.5 Contracts ................................................................................................................... 166

20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL .......................... 167 20.1 Hydrology, Hydrogeology and Water Management ............................................................ 167 20.1.1 Introduction ................................................................................................................ 167 20.1.2 Climate ...................................................................................................................... 167 20.1.3 Hydrology .................................................................................................................. 168 20.1.4 Hydrogeology ............................................................................................................ 169 20.1.5 Groundwater Flow at Cukaru Peki ............................................................................ 171 20.1.6 Water Supply Options ............................................................................................... 171 20.1.7 Mine Dewatering........................................................................................................ 174 20.1.8 Mining and Groundwater Impacts ............................................................................. 174 20.1.9 Conclusions ............................................................................................................... 175 20.1.10 Risks and Opportunities ....................................................................................... 175 20.2 Geochemistry....................................................................................................................... 175 20.2.1 Introduction ................................................................................................................ 175 20.2.2 Risk & Opportunities .................................................................................................. 176 20.3 Mine Waste Management: Tailings ..................................................................................... 176 20.3.1 Introduction ................................................................................................................ 176 20.3.2 Assumptions .............................................................................................................. 176 20.3.3 Site Selection Process .............................................................................................. 177 20.3.4 Design ....................................................................................................................... 178 20.3.5 Cost Assumptions...................................................................................................... 182 20.3.6 Scoping Level Cost Estimate .................................................................................... 183 20.3.7 Conclusions ............................................................................................................... 184 20.3.8 Risks and Opportunities ............................................................................................ 185 20.4 Environmental and Social Permitting and Management ..................................................... 185 20.4.1 Environmental and Social Setting ............................................................................. 185 20.4.2 Environmental and Social Baseline Studies .............................................................. 195 20.4.3 Environmental Approvals .......................................................................................... 196 20.4.4 Stakeholder Engagement .......................................................................................... 200 20.4.5 Key Technical Issues................................................................................................. 201 20.4.6 Closure Cost Estimate ............................................................................................... 202 20.4.7 Conclusion ................................................................................................................. 203

21 CAPITAL AND OPERATING COSTS ............................................................. 204 21.1 Scenario Summary .............................................................................................................. 204 21.2 Capital Costs ....................................................................................................................... 205 21.2.1 Basis of Estimate ....................................................................................................... 205 21.2.2 Mining Capital Costs.................................................................................................. 205 21.2.3 Processing ................................................................................................................. 206 21.2.4 Tailings Disposal ....................................................................................................... 207

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21.2.5 Other .......................................................................................................................... 208 21.2.6 Mine Closure Costs ................................................................................................... 208 21.3 Operating Costs ................................................................................................................... 208 21.3.1 Mining Operating Costs ............................................................................................. 208 21.3.2 Process and G&A ...................................................................................................... 210 21.4 Benchmarking ...................................................................................................................... 211 21.4.1 Introduction ................................................................................................................ 211 21.4.2 Capital Expenditures ................................................................................................. 211 21.4.3 Operating Expenditures ............................................................................................. 212

22 ECONOMIC ANALYSIS .................................................................................. 214 22.1 Introduction .......................................................................................................................... 214 22.2 Economic Assessment Process .......................................................................................... 214 22.2.1 General Assumptions ................................................................................................ 214 22.2.2 Commodity Price Assumptions ................................................................................. 215 22.3 Mine and Physical Assumptions .......................................................................................... 215 22.3.1 Mining ........................................................................................................................ 215 22.3.2 Process, Smelting and Refining ................................................................................ 217 22.4 Operating and Capital Costs ............................................................................................... 217 22.5 Revenue & Cash Flow Projections ...................................................................................... 218 22.6 Base Case Project Sensitivities ........................................................................................... 222 22.6.1 Single Parameter ....................................................................................................... 222 22.6.2 Twin Parameter ......................................................................................................... 222 22.7 Cash Costs .......................................................................................................................... 224

23 ADJACENT PROPERTIES ............................................................................. 225 24 OTHER RELEVANT DATA ............................................................................. 227 25 INTERPRETATIONS AND CONCLUSIONS ................................................... 228 25.1 Conclusions ......................................................................................................................... 228 25.2 Risks .................................................................................................................................... 230 25.3 Opportunities ....................................................................................................................... 230

26 RECOMMENDATIONS ................................................................................... 231 26.1 Introduction .......................................................................................................................... 231 26.2 Tasks required to meet the Development Schedule ........................................................... 231 26.3 Tasks required for a Pre-feasibility Study ............................................................................ 232 26.3.1 Mineral Resource Estimation .................................................................................... 232 26.3.2 Geotechnical .............................................................................................................. 233 26.3.3 Hydrology and Hydrogeological ................................................................................ 233 26.3.4 Mining ........................................................................................................................ 234 26.3.5 Infrastructure ............................................................................................................. 234 26.3.6 Metallurgical Testing.................................................................................................. 234 26.3.7 Mine Waste Disposal ................................................................................................. 236 U6782 Cukaru Peki PEA Report v30.docx

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26.3.8 Geochemistry ............................................................................................................ 236 26.3.9 Market Study ............................................................................................................. 237 26.3.10 Environmental and Social Studies ....................................................................... 237 26.3.11 Economics ............................................................................................................ 238 26.3.12 Project Development Schedule ............................................................................ 238

27 REFERENCES ................................................................................................ 240 28 CERTIFICATES OF QUALIFICATION ............................................................ 243

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List of Tables Table ES1-1: Table ES1-2: Table 4-1: Table 4-2: Table 4-3: Table 7-1: Table 10-1: Table 11-1: Table 12-1: Table 13-1: Table 13-2: Table 13-3: Table 13-4: Table 13-5: Table 13-6: Table 13-7: Table 13-8: Table 13-9: Table 13-10: Table 13-11: Table 13-12: Table 14-1: Table 14-2: Table 14-3: Table 14-4: Table 14-5: Table 14-6: Table 14-7: Table 14-8: Table 14-9: Table 14-10: Table 14-11: Table 16-1: Table 16-2: Table 16-3: Table 16-4: Table 16-5: Table 16-6: Table 16-7: Table 16-8: Table 16-9: Table 16-10: Table 16-11: Table 16-12: Table 16-13: Table 17-1: Table 19-1: Table 20-1: Table 20-2: Table 20-3: Table 20-4: Table 20-5:

SRK Mineral Resource Statement as at 31 March 2016 for the Cukaru Peki Deposit Upper Zone .................................................................................................................... 4 Mine Gate Production (Base Case)............................................................................... 5 Brestovac-Metovnica Exploration Permit .................................................................... 35 Brestovac-Zapad Exploration Permit........................................................................... 35 Mining Permits Required for the Project...................................................................... 38 Indicator minerals for alteration assemblages ............................................................. 56 Summary of Cukaru Peki Drilling as at 14 December 2015* ...................................... 68 Summary of density per mineralisation domain .......................................................... 73 Summary of Certified Reference Material for copper, gold and arsenic submitted by Rakita in sample submissions ..................................................................................... 78 Scope of Testwork ....................................................................................................... 83 Detailed Scope of Testwork ........................................................................................ 83 Head Characterisation of Composites......................................................................... 84 FMC Comminution Test Results ................................................................................. 85 Summary of XRD Results on Each Bulk Composite ................................................... 86 Rougher Flotation Summary ....................................................................................... 87 Cleaner Test (F8) Showing “Low arsenic” and “High arsenic” Concentrates.............. 88 Cleaner Test 12 Showing “Low arsenic” and “High arsenic” Concentrates ................ 88 Summary of Locked Cycle Tests (“Low arsenic” and “High arsenic” Concentrate Grades and Recoveries) ............................................................................................. 90 Bond Ball Mill Work Index Test Results ...................................................................... 92 DSO and Concentrate Multi-element assays .............................................................. 94 Principal Metallurgical Conclusions ............................................................................. 95 Raw length-weighted domain statistics for copper, gold and arsenic ....................... 100 Summary of Mineralisation Zones at the Cukaru Peki Project .................................. 102 Composite Statistics for Copper and Gold ................................................................ 104 Composite Statistics for Arsenic ................................................................................ 104 Summary of modelled semi-variogram parameters for the Cukaru Peki Mineralisation domain KZONE 300 .................................................................................................. 105 Details of Block Model Dimensions for Grade Estimation* ....................................... 106 Summary of Final Estimation Parameters for Cukaru Peki ....................................... 107 Summary Block Statistics for Ordinary Kriging and Inverse Distance Weighting Estimation Methods ................................................................................................... 111 SRK Mineral Resource Statement as at 31 March 2016 for the Cukaru Peki Deposit prepared in accordance with the CIM Code .............................................................. 113 Gradations for Indicated Material at Cukaru Peki at various %Cu Cut-off Grades ... 114 Gradations for Inferred Material at Cukaru Peki at various %Cu Cut-off Grades ..... 114 Geotechnical parameter ranges for worst, likely and best rock mass conditions in the DSO deposit .............................................................................................................. 117 Q’ and N Parameters for HGMS............................................................................... 117 Tonnes and grade for stope optimisation runs .......................................................... 127 Inputs and assumptions for Long Hole Open Stoping, stope optimisation process 9.3% diluted Cu% cut-off ........................................................................................... 129 Benefits and limitations of haulage options ............................................................... 130 Development parameters .......................................................................................... 134 Development requirements over life of project .......................................................... 135 Estimated Development cost..................................................................................... 136 Dilution and Recovery Parameters............................................................................ 141 Material available in cut off groups ............................................................................ 141 Production statistics for DSO and Flotation stages ................................................... 142 Mining Equipment ...................................................................................................... 145 Backfill statistics ........................................................................................................ 148 Mine Gate Production ................................................................................................ 155 Payment Terms used in the Technical Economic Model .......................................... 164 3 Monthly estimates of Brestovacka river flow (m /s) .................................................. 173 3 Catchment Yield for the Brestovak River (m /s) ........................................................ 173 3 Mean Surface Runoff for each Water Supply Dam option (m /s).............................. 174 Proposed Alternatives Advantages and Disadvantages of Site Location ................. 178 SLOS – Staged TSF Dam Construction Option 1 ..................................................... 179

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Table 20-6: Table 20-7: Table 20-8: Table 20-9: Table 20-10: Table 20-11: Table 20-12: Table 20-13: Table 20-14: Table 21-1: Table 21-2: Table 21-3: Table 21-4: Table 21-5: Table 21-6: Table 21-7: Table 21-8: Table 22-1: Table 22-2: Table 22-3: Table 22-4: Table 22-5: Table 22-6: Table 22-7: Table 22-8: Table 22-9:

SLOS – Staged TSF Dam Construction Option 2 ..................................................... 179 Capital costs for TSF Option 1 (Starter Dam, years 1 to 5) ...................................... 183 Operating and Sustaining Capital costs for TSF Option 1 ........................................ 183 Capital costs for TSF Option 2 (Starter Dam, years 1 to 5) ...................................... 184 Operating and Sustaining Capital costs for TSF Option 2 ........................................ 184 Environmental Baseline Studies undertaken for the Project in 2015 ........................ 195 Environmental Approvals Required for Mining and Building Permits........................ 196 Types of Environmental and Planning Approvals Required ...................................... 197 Steps in the EIA Approval Process Applied in Serbia ............................................... 199 Scenario Summary .................................................................................................... 204 Pre-production mining capital expenditure ................................................................ 206 Processing capital expenditure ................................................................................. 207 TSF Capital Expenditure ........................................................................................... 207 Breakdown of operating costs per mining phase ...................................................... 209 Breakdown of unit operating costs per mining phase ............................................... 209 Major consumable input prices.................................................................................. 210 Summary of production and capital expenditure assumptions ................................. 211 Mine Production Summary ........................................................................................ 215 DSO and Concentrate Production Summary Post Blending ..................................... 217 Smelter Terms ........................................................................................................... 217 Capital Cost Summary for Base Case including Extended Case ............................. 218 Operating Cost Summary for Base Case including Extended Case ......................... 218 TEM Summary........................................................................................................... 220 2 Parameter NPV Sensitivities .................................................................................. 223 2 Metal Price NPV Sensitivities ................................................................................. 223 Cash Costs USD / lb Cu ............................................................................................ 224

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List of Figures Figure 4-1: Figure 4-2: Figure 4-3: Figure 4-4: Figure 7-1: Figure 7-2: Figure 7-3: Figure 7-4: Figure 7-5: Figure 7-6: Figure 7-7: Figure 7-8: Figure 7-9: Figure 8-1: Figure 8-2: Figure 9-1: Figure 9-2: Figure 10-1: Figure 10-2: Figure 10-3: Figure 12-1: Figure 12-2: Figure 13-1: Figure 13-2: Figure 14-1: Figure 14-2: Figure 14-3: Figure 14-4: Figure 14-5: Figure 14-6: Figure 14-7: Figure 14-8: Figure 14-9: Figure 14-10: Figure 16-1: Figure 16-2: Figure 16-3: Figure 16-4: Figure 16-5: Figure 16-6: Figure 16-7: Figure 16-8:

Location Map of the Project (Source: SRK, modified from Esri, USGS, NOAA basemap, 2016)........................................................................................................... 29 Brestovac-Metovnica and Brestovac-Zapad Exploration Permits (Source: Resevoir, 2016) ........................................................................................................................... 34 Original Brestovac and Metovnica areas (Source: SRK, modified from Esri, USGS, NOAA basemap, 2014) ............................................................................................... 36 Project Development Schedule ................................................................................... 42 Late Cretaceous and Tertiary metallogenic belts in SE Europe, (modified from Sutphin et al, 2013) ................................................................................................................... 51 Geological map of the TMC, with Reservoir JV and 100% owned Permit areas (Source: Reservoir, modified from Banješević, 2010) ................................................. 53 Geology and exploration target areas in the Brestovac-Metovnica exploration permit (Source: Resevoir, 2016) ............................................................................................ 54 Typical Alteration Zonation section through Cukaru Peki UZ & LZ looking north ....... 57 Covellite-pyrite HSE mineralisation in advanced argillic altered andesite. FTMC1223, 698 m ........................................................................................................................... 58 Quartz-chalcopyrite-pyrite vein in potassic/phyllic altered andesite from the porphyry mineralised zone. FTMC1218, 1780 m ....................................................................... 58 High grade covellite breccia in massive pyrite. FTMC1223 480.9-487.7 m. ............... 59 Preliminary seismic interpretation illustrating numerous faults, looking NNW (Source: SRK, modified from Curtin university/ GEOING group doo, 2014) ............................. 60 Cross section looking northeast, illustrating the relationship between the plunging trend of elevated arsenic and key alteration features ................................................. 61 Plan and Cross-Section of the mineralisation in the Bor mining district (modified from Jankovic et al., 2002). Horizontal and vertical scales are the same. .......................... 63 Schematic NW-SE long section through the Lepanto enargite-Au deposit (from Hedenquist & Taran, 2013) ......................................................................................... 64 Line 60 in the CSAMT geophysical survey with annotation, looking northnorthwest (Source: Nikova, L., 2014) ........................................................................................... 66 Exploration work completed on the Brestovac-Metovnica exploration permit from 2006 to 2016 (Source: Reservoir, 2016) .............................................................................. 66 Location of new collars (red) completed since November 2013 NI43-101** .............. 68 Example cross section through the Cukaru Peki deposit ............................................ 70 Core Recovery at the Cukaru Peki Project ................................................................. 70 Cukaru Peki Standard Submissions versus Date ....................................................... 79 QAQC Standard Summary Charts from submission of Cukaru Peki Samples ........... 80 Theoretical copper grade versus recovery from mineralogy ....................................... 86 Locked Cycle Test flowsheet ....................................................................................... 89 Incremental and Log Histogram of Length Weighted Project Copper Assays ............ 98 Log histogram plots for copper and gold for the MPB domain .................................. 100 Schematic section of the Cukaru Peki deposit looking northeast ............................. 101 Cukaru Peki Mineralisation Model 3D View looking northwest (100 m slice) ........... 103 Log Histogram and Log Probability Plot for copper for the MPB domain at Cukaru Peki ............................................................................................................................ 104 Cukaru Peki Block Model Copper (%) Grade Distribution looking northnorthwest ... 108 Cukaru Peki Block Model Gold (g/t) Grade Distribution view NNW .......................... 109 Cukaru Peki Block Model Arsenic (%) Grade Distribution view NNW ...................... 109 Validation Plot (Easting) showing Block Model Estimates versus Sample Mean (25m Intervals) for MPB domain KZONE 102 for copper ................................................... 110 Plan view showing SRK’s wireframe-defined Mineral Resource Classification for the Cukaru Peki deposit view NNW ................................................................................ 112 Stability graph illustrating proposed stability zones (after Matthews et al. 1980, modified by Nickson 1992) ........................................................................................ 118 Geological wireframes for Zones 101, 102, 201 and 202 (looking NW) ................... 124 Geological wireframes including Zone 300 (lookingNW) .......................................... 124 Geological wireframes including Zone 400 (lookingNW) .......................................... 125 Distribution of stopes based on extraction sequence (lookingNW)........................... 126 DSO material with >9.3% diluted Cu and 9.3% diluted Cu and 6.0% Cu Float material

6.0% (in situ)

3,709

8.1

5.1

0.23

3

>2.5% Cu Float material

2.5% (in situ)

9,901

3.7

2.3

0.24

4

>1.0% Cu Float material

1.0% (in situ)

14,322

1.6

0.7

0.15

Run

Target Material

Cu % Cut-off

1

DSO material

2

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Figure 16-8:

Flotation material above 2.5% Cu (lookingNW)

Figure 16-9 :

All flotation material above 1% Cu (lookingNW)

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Table 16-4:

Inputs and assumptions for Long Hole optimisation process 9.3% diluted Cu% cut-off

Open

Stoping,

stope

Optimisation Run

Software

Geological Domain

In situ cutoff

Orientation

1A

Deswik.SO

101 & 102

6.0%

Along strike

30

15

30

LHOS Primary/Secondary Stoping

1B

Deswik.ASD

101 & 102

6.0%

Across strike

30

20

30

LHOS Tertiary Stoping

2A

Deswik.SO

101 & 102

6.0%

Along strike

30

15

30

LHOS Primary/Secondary Stoping

2B

Deswik.ASD

101 & 102

6.0%

Across strike

30

20

30

LHOS Tertiary Stoping

3

Deswik.SO

Depleted 300

2.5%

Along strike

30

15

30

LHOS Primary/Secondary Stoping

4

Deswik.SO

Depleted 300

1.0%

Along strike

30

15

30

LHOS Primary/Secondary Stoping

Block Size (m)

Mining Method

16.5 Access Options Three options for accessing the underground operation were considered. These were via production shaft, conveyor in decline and truck haulage in decline.

16.5.1 Production Shaft It is assumed that a 6.5 m shaft would be constructed to a depth of 700 m, allowing for access to the entire UZ operation. Surface infrastructure would be centred around the headframe reducing surface impact.

16.5.2 Conveyor in 1:5 decline A 1:5 decline would be constructed to target an area adjacent to the deposit. A second parallel decline would be constructed for equipment and personnel access. These declines would have 4.5 x 4.5 m profiles. An orepass and plate feeder arrangement would be constructed at the head of the conveyor, allowing for a minor surge capacity and continuous feed onto the conveyor. Surface infrastructure would include transfer towers and surface conveyors to deliver mined material to the processing facility.

16.5.3 Truck haulage in 1:7 decline A 1:7 decline would be constructed to target an area adjacent to the deposit and then extended as development of the lower levels of the deposit progressed. A second, smaller parallel decline would be constructed for ventilation and ancillary equipment access. The primary decline would have a 6.5 x 6.5 m profile while the secondary decline would have a 4.5 x 4.5 m profile. It is assumed that the truck fleet would deliver mined material directly to stockpiles at the processing plant, which would be located less than 1 km from the portal location.Table 16-5 outlines the benefits and limitations of each option. Based on the points identified, the decision was taken to use a 1:7 dual decline for truck haulage. This option provided the fastest means to access the stoping areas and could be completed for the lowest initial capital cost.

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A full trade-off study will need to be completed as part of the Feasibility Study once additional geotechnical information becomes available. Table 16-5:

Benefits and limitations of haulage options

Method

Pros

Cons

Shaft Access

•

• •

• •

Lower operational cost per tonne of ore Capacity can easily accommodate a 2 Mtpa production rate Surface infrastructure can be positioned in an optimal location in relation to the shaft

• •

•

1:5 Decline with conveyor

•

•

•

Development cost of decline may be cheaper due to fewer metres over horizontal distance (steeper grade) Similar development time to 1:7 decline due to less metres required (1,810 m vs 2,560 m), but dip decreases productivity and conveyor needs to be installed Less initial capital than shaft option

• • • • •

•

1: 7 Decline with truck haulage

• • • •

Lowest initial capital cost Fastest option to develop overall as no need for conveyor infrastructure. Allows for adequate ventilation even after stoping commences. Assuming decline is straight, the option exists to convert the ancillary decline into a conveyor decline at a later stage, providing greater flexibility.

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• • •

High initial capital cost Slow rate of mining to bore out to required diameter of 6.5 m. Expected construction time of 2.5 years not including any ground support, civils or headframe construction. In order to access the 1.0% cut-off material a 700 m shaft is required. Geotechnical considerations of the Miocene layer mean that the shaft location may not be optimal. Extra development or additional cost may be incurred in geotechnical rehabilitation. If it is determined that the shaft needs to be relocated then there would not only be additional development cost, but also an ongoing penalty to underground haulage. All equipment must be brought down by shaft and rebuilt underground. Ventilation shaft is still required Higher initial capital than decline with truck haulage due to surface and subsurface conveyor infrastructure. Steep toes difficult to drill and blast leading to uneven floor Shorter rounds required for same size jumbo to stay on gradient. Dual decline is required for ventilation and access by equipment, however this decline is too steep to use as a secondary means of haulage in the event of major conveyor downtime. Decline must be straight unless additional cost is incurred for installation of transfer towers.

Development time for a conveyor is similar to a 1:5 decline but both are shorter than sinking a production shaft. Additional material must be removed due to larger primary decline and longer length at shallower gradient. Higher operating haulage cost per tonne of material extracted

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16.5.4 Decline Location Two options for the location of the decline referred to as the northern and southern decline positions were assessed. Both the northern and southern portal locations (Figure 16-10 and Figure 16-11) were relocated to higher ground to prevent inrush of water during flood events. The relocation also provides more space for surface infrastructure such as stockpiles, rehandling areas and mine ventilation equipment. Designs and schedules were completed for both options. The southern decline requires an additional 200 m of development to reach the stoping area, and would delay the start of production, by six weeks.

Figure 16-10: Northern Decline layout relative to stoping areas (looking north)

Figure 16-11: Southern Decline layout relative to stoping areas (lookingSW)

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For this study, the northern option is adopted. This decision was based on the portal location in relation to the expected location of the processing facility; however, other items should also be considered before final design, such as: •

geotechnical conditions for both options;

•

ongoing operating costs due to surface haulage lengths;

•

land status and ownership;

•

environmental and social restrictions which may be applied due to the proximity of populated areas to either the portals, access roads or both; and

•

Government requirements for safe blasting radius for development of box cut and decline.

16.6 Mine Development 16.6.1 Decline Development The initial underground development for the project is assumed to be twin declines at a gradient of 1:7 from the northern portal location. It is assumed that over the first six months, the rate of development will be two cuts per day over the twin declines. This will be increased to three cuts per day after six months. The declines will be developed using both jumbos and bolting machines, with a capital allowance made for one spare machine to cover any breakdowns. A capital allowance of USD4M has been made for the initial portal construction, which includes surface excavation of box cuts, Armco tunnels and geotechnical support. A larger 6.5 x 6.5 m decline will be excavated to accommodate Caterpillar 775 rigid body haul trucks, which will run in convoy along this decline, providing the primary method of haulage from the mine. These trucks will be loaded from an orepass system, periodically extended to lower levels as the mine develops. A smaller 4.5 x 4.5 m decline will be used for access and egress of ancillary equipment and will provide the intake ventilation route for the first two years of development. The larger decline will act as the return airway due to the haulage occurring along this path. Both declines will intake ventilating air when the main ventilation raise becomes operational in the third year after start-up.

16.6.2 Mine Development Detailed layouts have been made to the -250 mRL. Due to the similar layout of all levels development requirements for levels below the -250 mRL have been based on the upper levels. Once a final cut-off grade has been determined, the lower levels will be designed. The current cut-off grade of 1.0% Cu has been shown to be economic for the lower levels of the mine, but may not be optimal. The cost model allows for the continuation of the orepass system and the 6.5 x 6.5 m decline to extend the reach of the primary haulage. The cost model also allows for a fresh air raise to surface which is extended with production down to the -400 mRL. This extension may not be necessary, depending on the final annual tonnage of the mine. Figure 16-12 shows the general layout of the mine development including the northern decline, return ventilation raise orepass system and level development.

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Figure 16-12: General layout of the upper level mine development (looking north) The following ancillary development included in the schedule and cost model: •

crib room;

•

staged pumping area and sumps;

•

magazine;

•

detonator magazine; and

•

underground workshop for all servicing of drills/loaders/integrated tool (“IT”) Carriers.

Table 16-6 outlines the development profile and maximum rates of production. Actual production rates in the schedule are governed by the equipment and number of headings available.

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Table 16-6: Description

Cukaru Peki PEA – Main Report

Development parameters Dimensions (m)

Section Type

Section

Length (m)

Rate (m/day)

Cross Cut

4.5l x 4.5w

Section

Arch 5.5 x 5.5

25

4

Decline Block

5.5l x 5.5w

Section

Arch 5.5 x 5.5

25

4

Decline Narrow

5.5l x 5.5w

Section

Arch 5.5 x 5.5

25

4

Decline Nth 1

6.5l x 6.5w

Section

Arch 5.5 x 5.5

3.8

7.6

Decline Nth 2

4.5l x 4.5w

Section

Arch 4.5 x 4.5

3.8

7.6

Decline Sth 1

6.5l x 6.5w

Section

Arch 5.5 x 5.5

3.8

7.6

Decline Sth 2

4.5l x 4.5w

Section

Arch 4.5 x 4.5

3.8

7.6

Footwall Drive

4.5l x 4.5w

Section

Arch 4.5 x 4.5

25

4

Drive Block

4.5l x 4.5w

Section

Arch 4.5 x 4.5

10

4

Drive Narrow

4.5l x 4.5w

Section

Arch 4.5 x 4.5

10

4

Pump House

5.5l x 5.5w

Section

Arch 4.5 x 4.5

25

4

Return Air Drive

4.5l x 4.5w

Section

Arch 4.5 x 4.5

25

4

Return Air Rise

6.5 dia

Circle

Circular 6.5

25

3

External Stockpile

5.5l x 5.5w

Section

Arch 4.5 x 4.5

3.8

3.8

Internal Stockpile

5.5l x 5.5w

Section

Arch 4.5 x 4.5

25

4

Sump

4.5l x 4.5w

Section

Arch 4.0 x 4.0

25

4

5 dia

Circle

Circular 5.0

3.8

4

OP Access

4.5l x 4.5w

Section

Arch 4.5 x 4.5

3.8

4

Access Small

4.5l x 4.5w

Section

Arch 4.5 x 4.5

3.8

4

Magazine

20l x 15w x 6h

Rectangle

Arch 15.0 x 6.0

3.8

1.9

Detonator Magazine

5.5l x 5.5w

Section

Arch 5.5 x 5.5

3.8

1.9

Access LG

5.5l x 5.5w

Section

Arch 5.5 x 5.5

3.8

4

Orepass

Workshop A

6.5l x 6.5w

Section

Arch 6.5 x 6.5

3.8

1.9

Workshop B

8.5l x 10w x 6h

Rectangle

Arch 10 x 6.0

3.8

1.9

5.5l x 5.5w

Section

Arch 5.5 x 5.5

3.8

1.9

Crib Room

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16.6.3 Development Schedule Table 16-7 summarises the development required during the life of the operation. Table 16-8 shows the estimated development costs for the key development types within the project. Figure 16-13 shows the Life of Project development schedule. It is assumed that development commences in October 2017 for the purpose of this schedule. Development of the declines and associated infrastructure will take approximately 21 months. Mineralised material will be available for extraction in Q3 2019. For the purposes of this study, it has been assumed that any development that is contained within the decline and has an operational life of greater than two years is considered capital. This includes the main ventilation rise, orepass, workshops, crib rooms and footwall drives. All other development is considered to be an operating cost. Table 16-7:

Development requirements over life of project

Development Type

Units

Value

Waste Development (not vertical) Production development (4.5 x 4.5 m)

(m)

19,544

Access development (inc 4.5 x 4.5 m decline)

(m)

8,457

Infrastructure development - (5.5 x 5.5 m)

(m)

840

Infrastructure development – decline (6.5 x 6.5 m)

(m)

5,406

Total Waste Development

(m)

34,246

(m)

8,444

Development in mineralisation (not vertical) Production development Access development

(m)

76

Total Development in mineralisation

(m)

8,520

Total Development (not vertical)

(m)

42,766

Raise bore 6.5 m

(m)

843

Raise bore 5.5 m

(m)

804

Drill and Blast

(m)

403

Total Vertical Waste Development

(m)

2,050

Vertical Waste Development

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Table 16-8:

Cukaru Peki PEA – Main Report

Estimated Development cost Pre-Production DSO Unit Cost (USD/m)

Total Qty (m)

DSO

Total Cost (USD’000)

Unit Cost (USD/m)

Total Qty (m)

Total Cost (USD’000)

3,972

91

363

Misc Capital Dev.

5,794

21

122

Decline 6.5 x 6.5

5,794

3,069

17,781

3,972

528

2,098

Decline 5.5 x 5.5

5,405

0

0

3,583

0

0

Decline 4.5 x 4.5

5,077

2,524

12,815

3,256

0

0

Vent Raise 6.5

10,613

36

379

8,792

561

4,933

Vent Raise 5.5

9,003

0

0

7,182

0

0

Ore Pass 6.5

6,947

42

295

5,125

127

653

Development 4.5 x 4.5

4,763

951

4,527

2,941

2,997

8,815

Development 4 x 4

4,763

35

167

2,941

20

59

Ore Development 4.5 x 4.5

4,763

256

1,218

2,941

2,137

6,286

6,934

37,304

6,461

23,207

Total

2.5% Float Unit Cost (USD/m)

Total Qty (m)

1% Float Total Cost (USD’000)

Unit Cost (USD/m)

Total Qty (m)

Total Cost (USD’000)

Misc Capital Dev.

3,135

0

0

3,216

0

0

Decline 6.5 x 6.5

3,135

618

1,936

3,216

1,125

3,618

Decline 5.5 x 5.5

2,747

793

2,179

2,827

0

0

Decline 4.5 x 4.5

2,419

0

0

2,500

0

0

Vent Raise 6.5

7,955

233

1,853

8,036

150

1,205

Vent Raise 5.5

6,345

0

0

6,426

804

5,166

Ore Pass 6.5

4,288

96

412

4,369

0

0

Development 4.5 x 4.5

2,104

11,331

23,845

2,185

8,395

18,346

Development 4 x 4

2,104

20

42

2,185

1,728

3,776

Ore Development 4.5 x 4.5

2,140

6,077

12,788

2,185

19,168

43,055

Total

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50

109

12,252

32,220

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Development Schedule)

Development Metres

6,000

4,000

2,000

-

Waste Production Drive - 4.5 x 4.5 Waste Access Drive - 4.5 x 4.5 Ore Access Drive - 4.5 x 4.5

Waste Infra Dev - 5.5 x 5.5 Waste Infra Dev - Decline 6.5 x 6.5 Ore Production Drive - 4.5 x 4.5

Figure 16-13: Life of Project development schedule

16.6.4 Ground Support Due to limited knowledge regarding the host rock material encountered by the declines and subsequent development, the following assumptions have been made: •

all intersections will be cable bolted with 6 m bolts;

•

resin or chemical bolts will be installed at a frequency of 12-19 bolts/metre depending on tunnel size;

•

all declines will be shotcreted with 100 mm thickness and floor to floor coverage;

•

all production drives will be shotcreted with 50 mm thickness and floor to floor coverage; and

•

raises are bolted/meshed or shotcreted deepening on access.

An allowance has also been made to cable bolt the backs of the top primary stopes to reduce the risk of any failures at the contact between the mineralisation and the host rock.

16.6.5 Dewatering Dewatering has been considered and allowed for through the inclusion of internal sumps and a main pumping station in the design and cost model. The main pumping station is located off the decline at the -40 mRL. This will be fed by staged sump pumps until the 1.0% cut-off zone is reached. At this point, a second pumping station at the -270 mRL will be installed. According to the preliminary models received by SRK from the Client, the estimated inflows of groundwater into the Phase I mine workings (down to -200 mRL) range from 62 to 178 L/s 3 3 (223 to 641 m /h) into the decline, and 73 to 125 L/s (263 to 450 m /h) into the DSO section of the mine. Estimated groundwater inflows into the bottom section of the mine during Phase II 3 are 163 L/s (587 m /h). In addition water from backfilling operations will also need to be pumped from the mine. A maximum backfill rate of 2,000 tpd has been assumed for this study. Excluding water loss from hydration of cement fill and evaporation a water inflow from backfill operations of 850,000 L/day or an instantaneous drainage rate of 9.8 L/s is estimated.

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For the purposes of this study, an installed pumping capacity of 385 L/s is used, which exceeds estimated maximum inflow rates by over 20%. Pumping would be provided by a series of Sundyme LMV333 186 kW pumps, each with a pumping capacity of 60 L/s. As these are modular pumps, they will be added as the mine increases in depth. Seven pumps have been accounted for in the cost model. Each production level and decline sump will have a designed sump equipped with a 30 kW submersible pump with a 145 L/s capacity.

16.6.6 Ventilation A detailed ventilation study has not been undertaken for this study, but will form part of any ongoing technical studies. The primary ventilation circuit has been designed with a capacity based upon the total number of operational equipment as per the mining schedule. At the peak period in Year 8, a required 3 airflow of 628 m /s is estimated. An upcasting ventilation raise to exhaust mine air has been designed adjacent to the main part of the mineralisation. The raise, approximately 420 m deep extending from surface (390 mRL) to -30 mRL level would be excavated with a 6.5 m diameter raise bore. The raise location passes through approximately 150 m of Miocene material, which typically consists of cemented sedimentary rock strata. It is also possible that the uppermost 10 m will need to be excavated due to the presence of unconsolidated material. Although the use of a large diameter raise is considered suitable, -intensive ground support including an allowance for 100 mm shotcreting and wall bolting are included in the cost model. There is an option to relocate the vent raise some 650 m to avoid the influence of the Miocene area, however this would increase development costs by USD2.24M and delay start of production by six months. Although this would likely reduce the need for intensive ground support within the raise, additional ground support would be required along the 650 m ventilation drive. It would also induce an additional loss of air/increase in power cost due to extra airway resistance and shock losses. It is recommended that a geotechnical hole be drilled at the proposed location to determine the quality of the material. If necessary, a trade-off study should then be undertaken between this location and others with less or no Miocene interaction. Before the upcasting ventilation raise is commissioned, the 4.5 x 4.5 m secondary decline will act as an air intake while the 6.5 x 6.5 m primary decline will act as a return airway. The haulage fleet will operate in the larger decline which will keep exhaust fumes in the return airway. As workings are deepened to extract the 1.0% cut off material located towards the base of the UZ, an additional fresh air intake ventilation raise to downcast fresh air from the surface may be required. This will provide ventilation for mining activities to the -400 mRL level as the mine progresses through the 1.0% cut off material. This option has been included in the cost model. A total of 40 axial auxiliary fans and starter boxes are required, based on the number of active operations headings and levels to meet the 2 Mtpa production schedule. Mine heating has been included in the cost model as, based on the average monthly temperatures, it will be required from December through to February, with a peak load in January. Figure 16-14 shows the key components of the ventilation system. U6782 Cukaru Peki PEA Report v30.docx

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Figure 16-14: Primary ventilation circuit (Green – fresh air intake, Red – Exhaust air) (Looking North)

16.7 Mine Extraction design and Sequencing 16.7.1 LoM Production Scenarios Two cases were developed, with each case considering a north and south decline access. Of these four schedules, two schedules were assessed in the cost model, and from this the final schedule was chosen for further analysis and reporting. The targets for scheduling are driven by the ability to feed the plant and the overall copper production. The two cases examined were: •

Base case mining to a 2.5% Cu cut-off with 2 Mtpa maximum production; and

•

Extended Case mining to a 1.0% Cu cut-off with 2 Mtpa maximum production.

In all cases, waste tonnages from development below the 250 mRL were accounted for in the cost model by estimating the average development required per level. The Extended case using the Northern decline is described in more detail below. Figure 16-15 shows the production profile for this case.

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1% CU Cutoff - 2Mtpa Maximum Float Feed 2,500,000

Tonnes

2,000,000

Total Ore Handled Total Waste Handled

1,500,000 1,000,000 500,000 -

Figure 16-15: The production profile for the chosen mining scenario

16.7.2 Dilution and Recovery Dilution and recovery were applied to each stope based on the optimisation method and extraction sequence. Two types of dilution have been considered. Internal Dilution: dilution relating to the contacts between the mineralisation and host rock to form a stope shape within the PEA mine plan. This represents the inherent inaccuracies in the geological interpretation and the creation of such stope shapes. External Dilution: dilution resulting from mining practices such as drilling/excavating into host rock and backfilled stopes. Primary stopes are considered to have 10% external dilution; however, this dilution is caused by material with grade as it comes from the secondary stopes. There are exceptions to this where the primary stopes are adjacent to the geological boundaries. The Deswik.SO module adds 1 m of internal dilution to both the hanging wall and footwall of all stopes that border the geological wireframes, accounting for dilution from the host rock which may occur here. The calculation of Cu grade to be used for primary stope dilution is defined as the average stope grade – 1.75% while the calculation of gold grade to be used for primary stope dilution is defined as the average stope grade of 2 g/t gold. Secondary stopes also have a 10% external dilution applied with grades of 0% and 0 g/t respectively, to reflect that most dilution will come from backfill material. Most of the secondary stopes will have good dilution control as the boundary between the consolidated fill material and rock is very easy to determine by the use of both CMS surveys and drill feedback (penetration rates/inspection of cuttings etc). Internal dilution of 1 m has also been added to secondary stopes adjacent to geological boundaries. All stopes created by the Deswik.ASD have been assigned an internal dilution of 1 m on both the hanging wall and footwall as they typically cover the width of the mineralisation. Their external dilution has been decreased to 5% to account for the increased internal dilution.

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The recovery of the stopes is assumed to be 90% of tonnage for all cases. This is to account for material left behind by remote bogging. Due to the size and layout of the stopes, 60% of extraction material is estimated to be free dig, and the remaining 40% will be through the use of remotely operated machines. As the stopes are then filled with consolidated material, it is assumed that the 10% of material left in the stope is not recoverable from the stopes below. All development has a 5% dilution factor added to account for overbreak, and a 3% recovery loss on stope drives. Table 16-9 summarises the dilution and recovery applied to the primary, secondary and tertiary stopes as well as the stope drives. Table 16-10 gives the breakdown of material in the different cut off groups after dilution and recovery has been applied. Note that these figures are based largely on Inferred Mineral Resources and a Pre-feasibility study has yet to be completed, therefore these numbers should not be interpreted to be Mineral Reserves. Figure 16-16 shows a heat map of the grade distribution within the stopes. The highest grade material is clearly visible towards the top of the deposit. Table 16-9: Stope Type Primary Stope

Secondary Stope

Tertiary Stope

Dilution and Recovery Parameters Dilution/Recovery

Value

External Dilution

10%

10% external dilution applied at average Cu stope grade -1.75% and average stope grade of 2 g/t gold.

Internal Dilution

1.0 m

Only applied to stopes adjacent to geological outlines

Recovery

90%

External Dilution

10%

10% external dilution applied at a grade of 0% as dilution will be from backfill material.

Internal dilution

1.0 m

Only applied to stopes adjacent to geological outlines

Recovery

90%

External Dilution

5%

Internal Dilution

1.0 m

Recovery

90%

Dilution

5%

Over break during excavation

Recovery

97%

Miscellaneous loss of mineralised material

Development

Table 16-10:

Comments

Applied to both hanging wall and footwall of each stope

Material available in cut off groups

Run

Target Material

Cu % Cut-off

Mined kt

Mined Cu%

Mined Au (g/t)

Mined As %

1

DSO

9.3% (diluted)

1,905

13.0

10.6

0.20

1

Stockpiled Float

6.0% Cu

6.0% (in situ)

3,490

7.9

4.8

0.23

3

Float >2.5% Cu

2.5% (in situ)

9,814

3.4

2.2

0.22

4

Float >1.0% Cu

1.0% (in situ)

14,179

1.5

0.7

0.14

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Figure 16-16: Grade distribution within stopes (looking north)

16.7.3 Production Schedule The annual production target tonnage for DSO material is 70,000 tonnes contained copper in the product. Using a diluted cut-off grade of 9.3% Cu it was determined that approximately 1.9 Mt of material at an average feed grade of 13.0% Cu will be able to deliver this target, yielding a total of 247,680 t of contained copper. This equates to approximately 545,000 t of DSO per year, for 3.5 years. In order to provide continuous production from lower grade material, the flotation plant must be commissioned 3.5 years after the initial DSO production commences. The float plant feed material available at a 1.0% Cu cut-off grade is 27.5 Mt at an average grade of 3.01% Cu, although initial grades are much higher and decrease over time. From this material 762,820 t of copper would be produced using current metallurgical assumptions. This allows the plant to operate for six years at 70,000 t of contained copper production followed by a further 10 years at a declining contained copper output as the feed grade drops. Table 16-11 shows the production statistics for the two processing phases, while Figure 16-17 shows the source of plant feed and contained copper in product over time. Note that these figures are based largely on Inferred Mineral Resources and a Pre-feasibility study has yet to be completed, therefore these values should not be interpreted to be Mineral Reserves. Table 16-11:

Production statistics for DSO and Flotation stages DSO

Flotation

Mineralisation (‘000 tonnes)

1,906

27,574

Cu (%)

13.0

3.0

Au (g/t)

10.6

4.1

As (%)

0.20

0.18

Contained Copper in product (t)

247,680

762,820

Operating life (years)

3.5

16

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2.5

Plant Feed Tonnes Vs Product Tonnes - @ 2mtpa

80,000 70,000

2.0

60,000

1.5

50,000

1.0

30,000

40,000 20,000

0.5

10,000 -

0.0

Flotation Ore

DSO Ore

Contained CU in concentrate (t)

Plant Feed Tonnes (Millions)

SRK Consulting

Contained Copper

Figure 16-17: Plant feed tonnes and resulting contained copper tonnes

16.8 Operating Strategy 16.8.1 Overall Extraction Strategy The proposed extraction strategy for the project is to first mine the high grade DSO material followed by the lower grade flotation feed material to produce a concentrate. All extraction is via a twin decline exiting to the north of the deposit. Mined material will be stockpiled on surface near the portal location for rehandling and transport to the processing facility or to market. Approximately 190,000 t of lower grade mineralisation to be processed eventually by flotation will be stockpiled during the DSO phase. The overarching aim for this study has been to minimise start-up capital requirements, and to use the cash resulting from sale of DSO material to provide capital funding for mining and processing of the lower grade flotation material.

16.8.2 Haulage Equipment The haulage fleet selected for the project is dependent on the final cut-off grade, the milling capacity and the final depth of mining. Using the current assumptions, three options were examined: •

AD 40 articulated trucks (40 t capacity);

•

AD 60 articulated trucks (60 t capacity); and

•

a combination of Caterpillar 775 rigid body trucks (65 t capacity) for long haul from the mine and AD 60 articulated trucks for internal haulage to orepasses.

The third option has been selected as the preferred option and has been used in the development of the design and cost model. AD 40 articulated trucks Previous studies had recommended the 40 t capacity trucks to achieve production rates of 500 ktpa; however, these are ruled out due to the increase in desired annual output to 2 Mtpa.

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AD 60 articulated trucks AD 60 trucks are all-wheel drive and can operate on the declines and in internal drives. They are more suited to an underground environment than the CAT 775 trucks, but have a lower maximum haulage speed on the declines. They also have a higher capital cost. Quotes obtained for this study suggest that they are 66% more expensive than CAT 775 trucks. In an operation with AD 60 trucks providing all of the haulage, there would be no rehandling of material and both mineralised material and waste could be hauled to surface in the same truck. These trucks can operate in a 5.5 x 5.5 m decline with services and will also operate in the smaller 4.5 x 4.5 m stope drives. Selected Option: CAT 775 trucks for primary haulage with AD 60 articulated trucks for internal haulage This option provides the greatest flexibility for the operation at a lower initial equipment capital cost. This case has been used as the basis for the cost model. CAT 775 rigid body trucks would be used for primary haulage from a feeder plate/orepass arrangement at the base of the decline to haul mineralised material to the processing facility. AD 60 articulated trucks would be used for waste haulage and to feed the orepass from lower levels. It is expected that there will be approximately 40% additional rehandle in this scenario. Direct feed by loaders from the stope to the orepass is assumed to be 60%. CAT 775 trucks require a minimum decline profile of 6.5 x 6.5 m to operate safely (including service installation). In this study, the dual decline has been designed along the main haulage way (for a distance of 2.5 km) which then joins the single 6.5 x 6.5 m decline. This approach allows for the relocation of the orepass/feeder plate and to extend the CAT 775 haul deeper into the operation as the working areas progress downward. Most operations taking this approach operate the haul trucks in convoy. This reduces the need to manoeuver or park the trucks in the decline to allow for passing. It also increases productivity through better utilisation of the truck fleet. Use of the CAT 775 trucks requires adequate drainage and continuous road maintenance within the decline, and this has been factored into the design and cost model. This option requires a greater number of trucks than the AD 60 articulated truck option, however the overall purchase price is lower, and the use of a plate feeder allows for internal stockpiling for up to three days, reducing the impact of production issues from individual working areas. This balances the production and haulage profiles and reduces stand-down of trucks when waste is being mined. The use of the orepass/plate feeder arrangement and convey haulage results in a highly productive haulage unit not usually associated with underground trucks. Use of the CAT 775 trucks for primary haulage and installing a 6.5 x 6.5 m primary decline allows the operation to increases production rates greater than 2 Mtpa with ease, simply by adding additional trucks. It is considered a better future proofing option than using AD 60 articulated trucks only.

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Cukaru Peki PEA – Main Report

16.8.3 Other Mining Equipment The Project includes other mining equipment, both primary and ancillary. Table 16-12 lists the type and number of additional units required during the life of the Project including estimated capital and operating costs. Replacement purchases have not been considered for this level of study. This will need to be incorporated into the Feasibility Study; however, the impact on the first five years of operation will be negligible. Table 16-12:

Mining Equipment

Type/Purpose

Equipment

Max Number of units

Capital cost (each) USD

Operating cost USD

LHD

Caterpillar 2900G

4

1,116,000

0.56/t

Truck

Caterpillar 775

5

992,866

1.20/t

Truck

AD 60

2

1,490,000

1.59/t

Face Drill

Sandvik DD422i

3

947,800

3.01/drm

Production Drill

Atlas Copco Simba 364

2

850,000

4.92/drm

Shotcreter

CSU 8000

1

884,350

27.00/h

Cement Carrier

Concrete Truck

1

150,000

42.00/h

Integrated Tool Carrier

Caterpillar IT28/926M

2

133,650

41.24/h

ANFO Loader

Normet Charmec 1610B

2

666,800

41.92/h

Grader

Caterpillar 120H

2

368,100

55.00/h

Auxiliary Loader

Caterpillar 1700G

2

761,400

86.59/h

Light Vehicle

Toyota Landcruiser 78 Series

21

42,000

25.07/h

Mechanical Scaler

Jama SBU 800 Scaler

1

766,350

59.64/h

Rock bolter

Sandvik DS411

2

1,196,700

33.51/h

Cable bolter

Sandvik DS520-TC

1

1,196,700

33.51/h

Service Vehicles

Flat bed or tank truck

2

326,000

42.92/h

16.8.4 Stope Drilling and Extraction The primary stopes have been designed on a 30 m sublevel interval with 15 m wide stopes. These are assumed to be drilled with 3.2 m burden and 3.5 m spacing based upon recommendations by Reservoir Minerals. These parameters will be reviewed following further geotechnical work. Figure 16-18 shows a typical ring layout.

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Figure 16-18: Typical ring layout for primary stopes The stopes are extracted using a primary/secondary method with a cemented backfill material applied. The maximum current stope length of 30 m, which each stope assumed to be backfilled before a raise is drilled for the next stope. Each level has up to six primary stopes of 30 m in length along strike. A barricade wall is constructed at the end of the first stope before the stope is filled and allowed to cure (see Section 16.8.5). Figure 16-19 shows a typical stope configuration for primary and secondary stopes for a single level. The cost model assumes that the brows are cable bolted; however, this may not be necessary pending the further geotechnical work. The next phase of geotechnical drilling may permit the increase of stope length, which would reduce the requirement for a raise at every 30 m.

Figure 16-19: Typical stope configuration on a single level (plan view)

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16.8.5 Backfill A comprehensive backfill study is required to be completed during the next phase of work. For scheduling and costing purposes, it has been assumed that the fill used in this study is a combination Cemented Aggregate Fill (“CAF”) and Hydraulic Sand Fill (“HSF”). This gives a combination of high strength required to undercut the primary stopes which is provided by the CAF and ease of fill and delivery provided by the HSF. Paste fill was considered but ruled out after consultation with Reservoir Minerals based on the following conditions: •

Limited tailings available for paste fill until the flotation plant is operational some 3.5 years after production commences.

•

Tailings will contain sulphides and may not be suitable for use in backfill operations.

•

Tailings may still have a high residual gold grade which may warrant retreatment at some future point.

•

No environmental study has been undertaken to determine if tailings are suitable for paste fill.

•

No external operation exists in the vicinity of the mine which is able to provide tailings for backfill.

•

Lower capital cost for setting up a HSF plant and CAF mixing plant than for a paste plant.

The following constraints were applied to the schedule and cost models: •

additional pumping capacity required for water seepage from the filled stopes;

•

operating costs are dependent on cement content (higher cement content increases operating cost);

•

source of sand/aggregate available to be trucked to site is within 10 km; and

•

a curing time of 28 days is assumed necessary/adequate.

The primary stopes are filled with 50% CAF material followed by 50% HSF, meaning that for a 30 m stope the fill is comprised of 15 m of CAF below 15 m of HSF. Secondary and tertiary stopes are filled with 100% HSF as there is no requirement for undercutting. It is assumed that all stopes are filled to 100% capacity. Table 16-13 shows the estimated quantities for the backfill operation.

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Table 16-13:

Backfill statistics Driver

Units

Total (‘000)

Stope Tonnes

(t)

29,467

Stope Volume

In situ density = 2.96

3

(m )

Primary Stope Fill – CAF

50% of primary stopes 24% of all stopes

(m )

Primary Stope Fill – HSF

50% of primary stopes 24% of all stopes

(m )

Secondary and Tertiary Stope Fill - HSF

52% of all stopes

(m )

CAF Tonnes

CAF density = 2.8

(t)

6,675

HSF Tonnes

HSF density = 2.0

(t)

15,128

Cement in CAF

5% cement in CAF

(t)

334

Cement in HSF

6% cement in HSF

9,951

3

2,384

3

3

2,384 5,180

(t)

908

Total Cement

(t)

1,241

Aggregate in CAF

(t)

6,342

Sand in HSF

(t)

12,798

CAF Composition and Strength The aggregate required for the CAF material is to have the following characteristics: •

Los Angeles Abrasion (LAA) Index < 20;

•

intact rock UCS > 70 MPa;

•

avoid high pyrite or pyrrhotite content;

•