Pre-Feasibility Study, Report Prepared for Amarillo Gold Corporation

Technical Update on the Posse Gold Project, Brazil, September 2018

Prepared for:

Amarillo Gold Corporation

Prepared by:

SRK Project Number: ARO001

Qualified Persons: Anthony Stepcich, FAusIMM(CP) Keith Whitehouse, MAusIMM CP(Geo)

Date of Report: 12 September 2018 Effective Date: 12 September 2018 SRK Consulting Page i

Date and Signature Page SRK Consulting (Australasia) Pty Ltd Level 1, 10 Richardson Street West Perth Western Australia, 6005 Australia

Technical Update on the Posse Gold Project, Brazil, September 2018

Project Manager: Anthony Stepcich, FAusIMM(CP) Date of Report: 12 September 2018 Effective Date: 12 September 2018

Mr Anthony Stepcich, FAusIMM(CP) Principal Consultant (Project Evaluations) SRK Consulting (Australasia) Pty Ltd Qualified Person (QP)

Mr Keith Whitehouse, MAusIMM CP(Geo) Director Australian Field Exploration Services Pty Ltd (AEFS) Qualified Person (QP)

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Important Notice This report was prepared as a National Instrument 43-101 Technical Report for Amarillo Gold Corporation (Amarillo) by SRK Consulting (Australasia) Pty Ltd (SRK). The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in SRK’s services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, including without limitation, those sources listed in Section 3 - Reliance on Other Experts, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by Amarillo subject to the terms and conditions of its contract with SRK and relevant securities legislation. The contract permits Amarillo to file this report as a Technical Report with Canadian securities regulatory authorities pursuant to National Instrument 43-101, Standards of Disclosure for Mineral Projects. Except for the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk. SRK accepts no responsibility with respect to the opinions of those experts listed in Section 3 - Reliance on Other Experts nor determinations made by the Company with respect its obligation to file this Technical Report, or subsequent technical reports, nor any determinations as to the materiality of a mineral project to Amarillo, and SRK is under no obligation to update this Technical Report, except as may be agreed to between Amarillo and SRK by contract from time to time. 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.

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Table of Contents

Important Notice ...... ii 1 Executive Summary ...... 1 1.1 Property Description and Ownership ...... 1 1.2 Geology and Mineralization ...... 1 1.3 Status of Exploration ...... 1 1.4 Mineral Resource Estimate ...... 2 1.5 Geotechnical Engineering ...... 2 1.6 Mining and Operations ...... 3 1.7 Recoveries ...... 3 1.8 Operating and Capital Costs ...... 4 1.8.1 Operating Costs ...... 4 1.8.2 Capital Costs ...... 4 1.9 Economic Analysis ...... 5 1.10 Mineral Reserve Estimate ...... 6 1.11 Comparison to Previous Studies ...... 7 1.12 Qualified Person’s Conclusions and Recommendations ...... 7 1.12.1 Conclusions ...... 7 1.12.2 Recommendations ...... 8 2 Introduction ...... 10 3 Reliance on Other Experts ...... 11 4 Property Description and Location ...... 12 5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ...... 16 5.1 Accessibility ...... 16 5.2 Climate ...... 16 5.3 Local Resources ...... 16 5.4 Infrastructure ...... 16 5.5 Physiography ...... 17 6 History ...... 18 7 Geological Setting and Mineralization ...... 20 7.1 Local Geology ...... 20 8 Deposit Types ...... 23 9 Exploration ...... 24 10 Drilling ...... 25 11 Sample Preparation, Analyses, and Security ...... 26 12 Data Verification ...... 27 13 Mineral Processing and Metallurgical Testing ...... 28

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14 Mineral Resource Estimates ...... 29 14.1 Historical Reports ...... 30 14.2 2018 Mineral Resource ...... 30 14.2.1 Data Used ...... 30 14.2.2 Wireframing ...... 31 14.2.3 Compositing and Statistics ...... 33 14.2.4 Variography ...... 36 14.2.5 Block Model ...... 40 14.2.6 Interpolation ...... 41 14.2.7 Model Classification and Reporting ...... 42 14.2.8 Mineral Resource ...... 44 14.3 Interpretation and Conclusions ...... 46 15 Mineral Reserve Estimates ...... 48 16 Mining Methods ...... 50 16.1 Geotechnical Design Parameters ...... 50 16.1.1 Footwall Amphibolite ...... 51 16.2 Mining ...... 54 16.2.1 Introduction ...... 54 16.2.2 Mining Model ...... 54 16.2.3 Pit Optimization ...... 54 16.2.4 Mine Layout ...... 55 16.2.5 Life of Mine Plan ...... 56 16.3 Production Scheduling ...... 59 16.3.1 Schedule Assumptions ...... 59 16.3.2 Schedule Results ...... 59 16.3.3 Mining Equipment and Operations ...... 66 16.3.4 Mining Labour Requirements ...... 67 17 Recovery Methods ...... 68 17.1 Variance of Grind Size ...... 68 18 Project Infrastructure ...... 69 19 Market Studies and Contracts ...... 72 20 Environmental Studies, Permitting, and Social or Community Impact ...... 73 20.1 Physical and Natural Environment ...... 73 20.2 Hydrology and Hydrogeology ...... 74 20.3 Social Environment ...... 74 20.4 Waste and Tailings Disposal, Monitoring and Water Management ...... 74 20.5 Permitting ...... 75 20.6 Mine Closure ...... 75 21 Capital and Operating Costs ...... 76 21.1 Approach ...... 76

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21.2 Operating Costs ...... 76 21.3 Capital Costs ...... 77 22 Economic Analysis ...... 80 22.1 Introduction ...... 80 22.2 Cashflow Assumptions ...... 80 22.2.1 Mine Production Sequence ...... 80 22.2.2 Operating and Capital Costs ...... 80 22.2.3 Metallurgical Recovery ...... 80 22.2.4 Metal Prices and Net Revenue ...... 80 22.2.5 Salvage Value ...... 81 22.2.6 Taxes ...... 81 22.2.7 Working Capital ...... 81 22.2.8 Reclamation ...... 81 22.3 Discounted Cashflow Result ...... 81 22.4 Sensitivity Analysis...... 82 22.5 Comparison to Previous Studies ...... 84 23 Adjacent Properties ...... 85 24 Other Relevant Data and Information ...... 86 25 Interpretation and Conclusions...... 87 26 Recommendations ...... 88 27 References ...... 90 28 Certificate of Qualified Person ...... 91 29 Certificate of Authorship ...... 92

List of Tables Table 1-1: 2018 Mineral Resource ...... 2 Table 1-2: Summary of LOM physicals ...... 3 Table 1-3: LOM cash cost estimate...... 4 Table 1-4: Summary of capital expenditure ...... 5 Table 1-5: LOM financials ...... 5 Table 1-6: DCF result ...... 6 Table 1-7: September 2018 Mineral Reserve estimate* ...... 6 Table 1-8: Comparison to previous studies ...... 7 Table 4-1: Concession and tenement schedule ...... 13 Table 8-1: Significant deposits in the Mara Rosa region...... 23 Table 14-1: Mineral Resource estimate - 2018 ...... 29 Table 14-2: Summary statistics for raw assays in the mineralised zone ...... 34 Table 14-3: Summary statistics for 1 m composite assays in the mineralised zone ...... 34 Table 14-4: Semi-variogram model parameters ...... 38

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Table 14-5: Percentile values used for MIK thresholds...... 38 Table 14-6: Bins and values used for MIK thresholds in the Main Hi zone...... 39 Table 14-7: Cross-validation results ...... 39 Table 14-8: Model dimensions ...... 40 Table 14-9: Model rotations ...... 40 Table 14-10: Model dimensions ...... 40 Table 14-11: Assigned density values...... 41 Table 14-12: Model parameters for interpolation runs other than Main Hi zone ...... 41 Table 14-13: Model parameters for interpolation runs for Main Hi zone ...... 41 Table 14-14: Search parameters for each interpolation run ...... 41 Table 14-15: Block classification scheme ...... 44 Table 14-16: Mineral Resource summary - 2018 ...... 45 Table 14-17: Mineral Resource summary 2018 - Measured and Indicated only ...... 45 Table 14-18: Comparison between 2011 Mineral Resource and 2018 Mineral Resource ...... 47 Table 15-1: September 2018 Mineral Reserve estimate* ...... 49 Table 16-1: Coffey geotechnical recommendations 2013 ...... 50 Table 16-2: SRK-optimized Eastern wall pit design parameters ...... 53 Table 16-3: Estimation of cut-off grade ...... 54 Table 16-4: Pit optimization parameters ...... 54 Table 16-5: Ramp design criteria ...... 59 Table 16-6: Mining physicals ...... 60 Table 16-7: Processing physicals ...... 63 Table 16-8: Mine equipment ...... 66 Table 16-9: Blasting parameters ...... 67 Table 18-1: Staged storage capacity, quantities and implementation ...... 69 Table 21-1: Mining variable operating costs ...... 76 Table 21-2: Estimate of fixed costs ...... 77 Table 21-3: Processing costs summary ...... 77 Table 21-4: LOM cash cost estimate...... 77 Table 21-5: Summary of capital expenditure ...... 78 Table 21-6: LOM operating and capital cost summary ...... 79 Table 22-1: Net revenue calculation...... 80 Table 22-2: LOM physicals ...... 81 Table 22-3: LOM financials ...... 81 Table 22-4: DCF result ...... 82 Table 22-5: Sensitivity analysis result ...... 82 Table 22-6: Sensitivity to Revenue...... 83 Table 22-7: Comparison to previous studies ...... 84

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List of Figures Figure 1-1: Sensitivity analysis ...... 6 Figure 4-1: Location of Amarillo’s Mara Rosa properties ...... 15 Figure 5-1: Mara Rosa and surrounding towns ...... 17 Figure 7-1: Geology of the Posse deposit ...... 22 Figure 9-1: Location of drill holes in the 2018 drill program ...... 24 Figure 13-1: Proposed flowchart for the operation ...... 28 Figure 14-1: Posse wireframes and drill holes looking north east ...... 32 Figure 14-2: Posse wireframes and drill holes looking north along the strike of the basalt dyke...... 32 Figure 14-3: Posse wireframes and drill holes looking down ...... 32 Figure 14-4: Surfaces used to control zonation in the block model ...... 33 Figure 14-5: Volume, in yellow between backfill and mined surface in relation to the orebody ...... 33 Figure 14-6: Grades of 1 m composites in the hanging Wall Zone ...... 35 Figure 14-7: Grades of 1 m composites in the Main Zone ...... 35 Figure 14-8: Grades of 1 m composites in the Footwall Zone ...... 36 Figure 14-9: Semi-variogram map of 1 m composites in the mineralization zone ...... 36 Figure 14-10: Directional semi-variograms for the three mineralization zones ...... 37 Figure 14-11: Directional semi-variograms for the short range high grade portion of the Main zone ...... 37 Figure 14-12: QQ plot of the modelled data vs input data...... 42 Figure 14-13: Ore classification boundaries used in the model ...... 43 Figure 14-14: Grade-tonnage curve ...... 46 Figure 15-1: LOM variable cut-off grade ...... 48 Figure 16-1: Current geotechnical drilling with structural ATV disks shown ...... 50 Figure 16-2: Stereonet displaying structures below the footwall with strong NW dipping foliation, and dip variation for foliation presented in histogram ...... 51 Figure 16-3: East wall cross sections used for slope design ...... 52 Figure 16-4: Example of orebody dip variability, addressed by developing different FW blocks for slope design ...... 52 Figure 16-5: Proposed mine layout ...... 56 Figure 16-6: Final pit design and orebody (various views) ...... 57 Figure 16-7: Pit stages and final pit design ...... 58 Figure 16-8: Mining physicals ...... 62 Figure 16-9: Processing physicals ...... 64 Figure 16-10: Processing by grind size ...... 65 Figure 18-1: General layout of the TSF ...... 70 Figure 22-1: LOM cashflow and margin ...... 82 Figure 22-2: Sensitivity spider chart ...... 83

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List of Abbreviations

Abbreviation Meaning $ dollar sign % percentage sign °C Celsius sign µm micrometer 3D three dimensional ACME Acme Analytical Labs Ltd AEFS Australian Exploration Field Services Pty Ltd Amarillo Amarillo Gold Corporation ANFO Ammonium Nitrate and Fuel Oil (FO) ASL above sea level ATV acoustic televiewer Au gold symbol Barrick Barrick do Brasil bcm bank cubic metre BD Bulk Density BFA Bench Face Angle BH Bench height BHP now BHP Billiton Limited BoD Basis of Design BSA Bench Stack Angle BSH Bench stack height BVP BVP Engenharia C cohesion CAPEX capital expenditure CC Correlation Coefficient CCIC Caracle Creek International Consulting Inc CELG Brazil-based company engaged in the electric power industry CIL Carbon-in-Leach CIM Canadian Institute of Mining CTD conventional tailings disposal DFS Definitive Feasibility Study dmt Dry metric tonne DTM digital terrain model DXF Drawing Exchange Format E East E–W east-west FEL Front end loader FoS Factor of Safety

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Abbreviation Meaning FS Feasibility Study FSL full supply level FW Footwall g gram g/cc Gram per cubic centimeter, a density measure, see Kg/m3 g/t Gram per tonne GB geotechnical berm GBI geotechnical blockiness index GDM geotechnical domain model GIS geographic information system GMA Goiás Magmatic Arc GO Goiás state, Brazil GPS Global Positioning System GSBW geotechnical safety berm width GTR Grind-Throughput-Recovery ha hectares HCS Hoogvliet Contract Services HME heavy mining equipment Hr hydraulic radius HW Hanging Wall ICP Inductively Coupled Plasma IDS Inverse Distance Squared. A modelling algorithm where points are weighted by the square of the distance IOSA Indicative Overall Slope Angles IPC in-pit crushing IRA inter-ramp angle JORC Code Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves prepared by the Joint Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia (JORC), 2012. k thousand kg kilogram Kg/m3 Kilograms per cubic metre, the SI unit for measuring density, a density of 1000 kg/m3 is equivalent to 1 g/cc or 1 t/m3 or 1t/bcm kJ kilojoule kl kilolitre kt kilotonne kV kilovolt kW kilowatt l litre l/s Litres per second lcm Loose cubic metre

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Abbreviation Meaning LG Lerchs-Grossman LOM Life of Mine LOSA limiting overall slope angle m metre M million m RL metres reduced level m3 cubic metre MBL Metallica Brazil Ltda MCAF Mining Cost Adjustment Factor mE metres east MIK Multiple Indicator Kriging Minere Minere Mineração Ltda. MIP Maximum Intensity Projection MJSA Mineração Jenipapo S.A. ml millilitre Ml megalitgre mN metres north MR files Mara Rosa files MRMR Mining Rock Mass Rating mS metres south MSSO MineSight Scheduling Optimiser Mt million tonnes Mtpa million tonnes per annum MVA Mega Volt Amperes MW megawatt N north NE northeast NPV net present value NSR Net Smelter Return NW northwest OB Overburden OPEX operating expenditure OS Oversize oz Troy ounces; 31.103477 grams P80 Percent passing through specified mesh size PCAF Processing Cost Adjustment Factor PFS Pre-Feasibility Study PMF Probable Maximum Flood Prober Enterprise Optimization Software, Whittle Consulting Pty Ltd Q Barton Q value

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Abbreviation Meaning Q’ modified Q value QP Qualified Person QQ Quantile / Quantile plot RC reverse circulation RL reduced level RMR Rock Mass Rating ROM Run of Mine RQD Rock Quality Designation S South SBW Spill berm width SE southeast SG specific gravity SRK SRK Consulting (Australasia) Pty Ltd SRKBR SRK Consultores do Brasil Ltda SW southwest t tonne t/m3 Tonnes per cubic metre, a density measure, see Kg/m3 t/bcm Tonnes per bcm, a density measure, see Kg/m3 tpa tonnes per annum TSF Tailings Storage Facility TTD Thickened Tailings Disposal US United States UTM Universal Transverse Mercator W West WFS waste storage facility WMC Western Mining Corporation wmt wet metric tonne WSF Water Storage Facility

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1 Executive Summary A Pre-Feasibility Study (PFS) level open pit mining study has been completed on the Posse gold deposit owned by Amarillo Gold Corporation. The study consisted of the adjustment of the Mineral Resource Model, an Enterprise Optimization undertaken by Whittle Consulting, production scheduling, mining equipment and labour estimation, mining operating strategy, and mining cost adjustments from the previous PFS update. Capital and Operating costs have been estimated to within a precision range of +/-25%. No underground mining methods have been evaluated in this case.

This Executive Summary highlights the work undertaken in this September 2018 study and outlines the material changes between this study and previous studies.

1.1 Property Description and Ownership The Mara Rosa Property (also generally known and referred to as the Posse deposit, or the Project) is located in Goiás state (GO), central Brazil, approximately 6 km north of the town of Mara Rosa. The Project encompasses a land area of 2,552 ha across three mining concessions plus numerous exploration leases in areas surrounding the Posse mine area.

Amarillo Gold Corporation (Amarillo) visited the project in August 2003 and in October 2003 signed a letter of intent with Metallica Brazil Ltda (MBL) to purchase MBL and 100% of the Mara Rosa project. Amarillo is currently owner of 100% of the Posse Project.

1.2 Geology and Mineralization Amarillo’s land position within the Mara Rosa District primarily covers the Eastern Belt greenstone assemblage, with some coverage of the Western and Central belts as well. The Eastern Belt has a maximum thickness of 6 km, generally strikes to the northeast and dips moderately to steeply to the northwest.

The Posse deposit occurs in a regional thrust that probably acted as one of the primary dewatering conduits during the Neo-Proterozoic Brasiliano orogeny. The geophysical, geological and geochemical data available demonstrate that the Posse deposit occurs within a 50 km long shear zone with potassium alteration and lower order gold-copper-molybdenum mineralization. The Posse deposit has grey gneiss in the hanging wall of the fault and amphibolite, ‘greenstone’ in the footwall. Shearing of the Grey Gneiss has resulted in the formation of a distinct lithologic unit, a quartz-feldspar-mica schist (Posse Schist) that is characteristic of the Posse ore zone. This unit has been identified in several other areas, including the Posse footwall and on strike extensions of the Posse Ore Zone to the northeast. Shearing is most intense in the footwall.

The mineralization envelope at Posse is about 30 m thick and over 1 km long. It has a mylonitic appearance which is most noticeable in the footwall where shearing is the most intense. Higher intensity of shearing is associated with increased sulphide mineralization (up to about 4%), and a slight increase in metamorphic grade from greenschist to high greenschist facies in the hanging wall through to high greenschist/ low amphibolite facies in the footwall (biotite flakes and garnet alteration). Higher gold values are associated with increasing intensity of shearing and higher levels of sulphide mineralization.

1.3 Status of Exploration During the period from late 2012 until June 2018, no drilling was carried out or samples submitted for assay. Amarillo started a 56-hole drilling program at the Project in June 2018. The program consists of 30 reverse circulation (RC) drill holes, 18PRC001–18PRC030, for a total planned length of 3,150 m

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and 26 diamond drill holes, 18P047–18P072, with a total planned length of 7,345 m. At the time this Report was compiled, 14 of the RC holes and 15 of the diamond holes had been completed. Work is currently underway on four of the diamond holes. The drill program is expected to be completed towards the end of 2018.

1.4 Mineral Resource Estimate A Mineral Resource can only be declared for material which is considered to have reasonable prospects for economic extraction at some point in the future. The cut-off at which a resource is reported should also meet this criterion, it should not include material which does not have reasonable prospects to be mined and processed. The definition on a Mineral Reserve applies a specific set of economic parameters to a Mineral Resource to determine which portions of the Mineral Resource can be mined economically. In the case of the Posse deposit, economic modelling of the blocks in the model indicated that there are blocks with grades above 0.216 g/t Au that will be economic to mine. On this basis, the cut-off grade for the Mineral Resource has been set at 0.2 g/t Au. The Mineral Resource above a cut-off of 0.2 g/t Au declared for the Posse deposit is summarized in Table 1-1.

Table 1-1: 2018 Mineral Resource

Volume Tonnes Grade Metal % Category Above Domain Category (Mm3) (millions) (g/t Au) (koz) Metal 0.20 1.6 4.4 0.69 97 17 Measured 0.20 2.4 6.4 0.80 170 23 HW Indicated 0.20 1.3 3.6 0.53 62 19 Inferred 0.20 2.4 6.6 2.1 440 79 Measured 0.20 3.6 9.8 1.6 500 71 MAIN Indicated 0.20 2.2 6.0 1.3 250 75 Inferred 0.20 0.50 1.3 0.40 17 3 Measured 0.20 0.88 2.4 0.52 40 6 FW Indicated 0.20 0.58 1.6 0.39 20 6 Inferred 0.20 4.6 12 1.4 560 100 Measured 0.20 6.8 19 1.2 710 100 ALL Indicated 0.20 4.1 11 0.92 330 100 Inferred Notes: HW = Hanging Wall MAIN = Main FW = Footwall. All figures have been rounded to two significant figures. A cut-off grade of 0.2 g/t Au has been used. Due to rounding, numbers may not sum correctly.

The authors are not aware of any modifying factors such as mining methods, metallurgy, environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other factors which will have an impact on the Mineral Resource.

1.5 Geotechnical Engineering SRK reviewed the ‘DFS Geotechnical Assessment 2013’ previously prepared by Coffey. In this PFS update, SRK accepted the Coffey recommendations for the hanging wall gneiss, and the Northern and Southern schists; however, SRK has redesigned the Eastern Footwall in the amphibolite zone of the pit.

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In the Eastern Footwall, SRK adopted a variable inter-ramp angle (IRA) model to match the orebody plunge, represented by eight blocks. SRK performed kinematic stability analysis for the range of plunge within each domain to estimate the likely factor of safety (FoS) and potential bench failure volume. Empirical analyses were conducted to confirm the spill berm widths required to adequately catch any failed material, the results of which showed wider berms with shallower bench face angles are required. The variable pit geometry elements are defined by a wireframe which was used for design and optimization purposes.

1.6 Mining and Operations The Mineral Resource Model supplied by AEFS was converted to a Mineral Reserve Model using the following parameters:

• Ore loss of 3% • Ore dilution of 3%, dilutive material grading 0.16 g/t Au • Moisture content of 3% • A variable cut-off grade was used dependant on Mill grind size:

− A P80 45 µm grind size which had a higher recovery of 92.0%. The cut-off grade calculated for this grind size was 0.329 g/t Au

− A P80 75 µm grind size which had a lower recovery of 86.34%. The cut-off grade calculated for this grind size was 0.216 g/t Au • The Reserve Model assumes highly selective mining of the mineralized zones.

The objective of the Posse production scheduling was to meet the production needs of the mill and optimize net present value (NPV) while maintaining good operational practices and a safely operated mine.

Production scheduling involved definition of mining phases, and development of the ore stockpile strategy and mining schedule.

Table 1-2: Summary of LOM physicals

LOM physicals Units Value Material movement LOM Total material movement (TMM) Mt 139.2 Total waste movement Mt 115.3 Total ore mined Mt 23.8 Average ore grade g/t 1.42 Processing LOM Ore processed Mt 23.8 Ore grade processed g/t 1.42 Average recovery % 90.6% Gold produced koz 985

The LOM plan has an average annual production estimated at 143.8 koz Au over the first four years, and an average LOM production of 123.1 koz per year over eight years, and a total recovered gold production of 984.8 koz.

1.7 Recoveries In the Enterprise Optimization undertaken by Whittle Consulting, a strategy of grind size variation was implemented.

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The ultimate grind size selected for the mill affects the mill throughput, power consumption and gold recovery. Two Grind-Throughput-Recovery (GTR) scenarios have been modelled in the Techno-Economic model constructed for the Project.

The Enterprise Optimization has optimised the GTR of the mill over the LOM:

• The P80 45 µm grind size is used for the high-grade component of the ore, with the aim of maximising the recovery and gold production

• The P80 75 µm grind size is used for the lower grade portion of the orebody with the aim of maximising mill throughput, while accepting a lower recovery from the lower grade ore.

Amarillo is conducting ongoing Bond Work Index (BWI) testwork that will give guidance on the grind power demand.

1.8 Operating and Capital Costs

1.8.1 Operating Costs Unit operating costs used in this study are outlined below:

• An overall mining and haulage cost of US$2.04/t Total Material Mined (TMM) • A processing cost of US$11.78/t for the 45 µm grind size and a recovery of 92% at a cut-off grade of 0.329 g/t Au • A processing cost of US$7.25/t for the 75 µm grind size and a recovery of 86.34% at a cut-off grade of 0.216 g/t Au • An estimated selling cost of US$13.60/oz Au • Royalties of 6% of revenue • A gold price of US$1,300/oz • A power cost of US$0.08/kWh • A diesel price of US$0.90 per litre.

Table 1-3 shows the LOM Cash Cost estimate based on production of 985 koz Au.

Table 1-3: LOM cash cost estimate

LOM Cash Cost Estimate Total Cost (US$M) Unit Cost (US$/oz) Operating Costs Mining & Processing - Variable $417.8 $424.3 Mining & Processing - Fixed $73.7 $74.8 Site G&A $45.4 $46.1 Operating Cost $536.9 $545.2 Adjusted Operating Costs Refining, Transportation, Insurance $13.4 $13.6 Royalties $73.5 $74.7 Adjusted Operating Costs $623.8 $633.5 All-in Sustaining Cost (AISC) Sustaining capital $17.4 $17.7 Reclamation costs $4.0 $4.1 AISC $645.2 $655.3

1.8.2 Capital Costs The total forecast capital expenditure over the LOM is US$140.3M.

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This consists of US$122.9M of pre-production capital to bring the mine into operation, and US$17.4M of sustaining capital over the LOM. Table 1-4 summarises the Project capital expenditure.

Table 1-4: Summary of capital expenditure

PFS Case Parameters Basis (US$M) Mine Contractor $5.3 SRK - PFS Processing Capital $104.2 Amarillo-adj by exchange rate Tailings Dam $23.0 SRK - PFS Working Capital $7.7 SRK - PFS Total Capital $140.3

Salvage - contractor -$12.1 SRK - PFS Reclamation $4.0 SRK - PFS Total Capital - Salvage + Reclamation $132.2

1.9 Economic Analysis The economic analysis undertaken contains forward-looking information with regards to the Mineral Reserve estimates, commodity prices, exchange rates, proposed mine production plan, projected recovery rates and processing costs, infrastructure construction costs and schedule. The results of the economic analysis are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.

• SRK’s discounted cashflow (DCF) economic analysis had the following basis: − A Base Case gold price of US$1,300/oz was used − The analysis was based on 100% equity financing with no debt component − All revenues and costs are reported in ‘real’ constant US$ terms without escalation − SRK’s economic analysis is for the purpose of Mineral Reserve estimates only.

The LOM financials from the financial model are shown in Table 1-5. The resultant internal rate of return (IRR) is 50.8%.

Table 1-5: LOM financials

Total Financials LOM (US$M) Total net gold sales 1,193.3 Total variable costs 417.8 Total fixed costs 119.0 Total capital (including reclamation) 144.3 Tax (34%) 187.7 Net cashflow (undiscounted) 324.5 NPV @ 5% (post-tax) 244.3

Table 1-6 shows the results of the DCF analysis. The IRR produced is 50.8%.

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Table 1-6: DCF result

Total Result (US$M) NPV (post-tax @ 5% discounted rate) 244.3 NPV (post-tax @ 7.5% discounted rate) 212.7 NPV (post-tax @ 10% discounted rate) 185.4

Figure 1-1 shows the results of the sensitivity analysis.

Sensitivity $450.0 $400.0 $350.0 $300.0 $250.0 $200.0 $150.0 $100.0 $50.0 $0.0 -30% -20% -10% 0% 10% 20% 30%

OPEX NPV US$M @ 5% CAPEX NPV US$M @ 5% Revenue NPV US$M @ 5%

Figure 1-1: Sensitivity analysis

1.10 Mineral Reserve Estimate The Mineral Reserve estimate is presented Table 1-7.

Table 1-7: September 2018 Mineral Reserve estimate*

Note: *Open pit Mineral Reserves are reported at a variable cut-off grade dependent on GTR, assuming a metal price of US$1,300/oz, mining cost of US$2.04/t TMM, a variable processing cost dependent upon cut-off grade and GTR of Mill, an estimated selling cost of US$13.60/oz, a royalty of 6% of sales, and a variable processing recovery dependent on GTR. The Mineral Resources Reported are inclusive of the reported Mineral Reserve.

The Qualified Person responsible for the reporting of this Mineral Reserve is Anthony Stepcich, a Principal Consultant and full-time employee of SRK Consulting (Australasia) Pty Ltd. In Mr Stepcich’s opinion, the technical parameters that form the basis of this Mineral Reserve statement are reasonable.

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Other than discussed herein this report, Mr Stepcich is not aware of any mining, metallurgical, infrastructure, permitting, environmental, legal, title, taxation, socio-economic, marketing or other relevant factors that could materially affect the Mineral Reserve estimate.

1.11 Comparison to Previous Studies Table 1-8 compares the 2018 PFS Update by SRK with the two previous PFS reports, Coffey (2011) and SRK Consultores do Brasil Ltda (SRKBR) (2017).

Table 1-8: Comparison to previous studies

2011 Coffey PFS 2017 SRKBR PFS 2018 SRK PFS Category Units (Owner Operator) (Contractor Mine) (Contractor Mine) Exchange rate US$/BRL 1.9 3.2 3.6 Initial capital US$M 183.6 132.3 122.9 Sustaining capital US$M 11.4 16.5 17.4 Total LOM capital US$M 195 148.8 140.3 Post-tax NPV @5% US$M 178.5 178.3 244.3 Post-tax IRR % 26.6 35.2 50.8 Cash operating cost (excluding US$/oz 464 545 545 royalty & refining) Cash operating cost (including US$/oz 524 603 633 royalty & refining) AISC (including sustaining capital & US$/oz 546 627 655 closure) Tonnes of ore processed Mt 17.1 19 23.8 Grade of ore processed g/t 1.72 1.63 1.42 LOM strip ratio (waste: ore) t:t 8.2: 1 4.5: 1 4.84: 1 Resources Measured & Indicated koz 1,174 1,260 1,300 Resources cut-off grade g/t 0.50 0.35 0.20 Reserves Proven & Probable koz 945 998 1,087 Reserves cut-off grade g/t 0.5 0.38 Variable US$1,300/oz 0.85 US$1,100/oz Pit US$1,200/oz Pit Revenue Factor Shell Shell Gold price US$/oz Pit Shell US$1,200/oz US$1,200/oz US$1,300/oz Financials Financials Financials 3% loss & 3% 3% loss & 3% 3% loss & 3% Mining loss & dilution % dilution dilution dilution Metallurgical recovery % 92 92 Variable

1.12 Qualified Person’s Conclusions and Recommendations

1.12.1 Conclusions The economic analysis presented undertaken contains forward-looking information with regards to the Mineral Reserve estimates, commodity prices, exchange rates, proposed mine production plan, projected recovery rates and processing costs, infrastructure construction costs and schedule. The results of the economic analysis and are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.

SRK’s economic analysis is for the purpose for Mineral Reserve estimates only.

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The economic analysis performed from the results of the PFS demonstrates an economic viability of the Posse project using the assumptions considered. Therefore, in SRK’s opinion, the Mineral Reserves declared are valid.

On the basis of SRK’s analysis, the Posse project generates a post-tax IRR of 50.8%, a post-tax net present value (NPV5) of US$244.3M, and post-tax project payback of 2.6 years based on a gold price of US$1,300/oz and a US$/BRL exchange rate of 3.60.

SRK undertook a sensitivity analysis of the PFS economic analysis result. The Project’s economics are most sensitive to revenue and least sensitive to capital expenditure.

Based on the technical studies presented and the positive economics, SRK considers that development of the Project from PFS to a Feasibility Study (FS) level is warranted.

1.12.2 Recommendations Further works include the following:

• It is recommended that a further round of resource drilling be undertaken to convert material currently classified as Inferred Mineral Resource into Indicated Mineral Resource or a higher resource classification. • There are a number of geotechnical holes which have not been assayed. These should be sampled and assayed. • Further work should be under taken to define higher grade shoots in the orebody. These are likely to have a significant effect on the economics of the deposit. • With further drilling the resource model will need to be updated. This update should include the use of smaller blocks (5 m × 5 m × 5 m) to provide better granularity of the resource. As part of an updated Mineral Resource, the constraining wireframes used should be reviewed and revised where appropriate. Improvements to the understanding of the oxidation zones in the deposit would be very useful as they are likely to affect processing costs and recoveries. SRK notes that the smaller block size will need to be geostatistically supported by the drill spacing. • If there is a reasonable spread of assay results for elements other than gold, these should, after appropriate validation, be modelled especially if links can be made between associated mineralization and processing cost and recoveries. Similarly, the density (specific gravity, SG) should be included in the model as an estimated value, rather than a series of assigned values based on oxidation zone. • Further work to define the nature and extent of backfill in the historical WMC pits will allow better definition of the waste/ resource boundary. Such work may entail a LiDAR survey and a bathymetric survey in the old pits. • It is further suggested that when the resource is remodelled, both inverse distance squared (IDS) and Ordinary Kriging (OK) interpolations should be carried out and compared to the Multiple Indicator Kriging (MIK) method with the modelling process peer reviewed prior to publication. • For the PFS update, SRK has assumed that the footwall orebody plunge and foliation are coincident (the regional geological data suggests foliation dips in the 35°– 55° range which would support this assumption); at the FS stage, it may be possible to refine this if more structural data is available. • There is ~30 m of oxide on the footwall side which will need to be checked in the FS design – it is expected that horizontal drains and berm drainage will be needed to prevent zones of erosion coalescing to create larger zones of failure and pit seepage which will result in higher water

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management costs. This design can be reassessed when accessing the old pits to identify suitable geometry. • Major structures need to be mapped in the old pit once access is re-established and used to develop a working structural geological model to assist pit design. • Standard ground control/ slope management procedures need to be adopted so that the design assumptions are validated during mining and the design is further optimized. Mapping of the footwall structures will be very important to maintain the optimal pit production as well as checking for the potential for adverse footwall structures which could be unstable. • Good quality blasting of final walls and major intermediate cutbacks will be critical to good performance; pre-splitting (or similar blasting techniques) should be adopted. • Amarillo is conducting ongoing Bond Work Index (BWI) testwork that will give guidance on the grind power demand. • The power demand for the nominal grind is a key element in the GTR model. Ongoing testwork is recommended. • SRK recommends that during any future Definitive Feasibility Study (DFS), a complete review of all operating and capital costs assumptions will be necessary to ensure a DFS level of estimation accuracy. All major equipment should be re-quoted and new contractor quotes obtained for any future DFS process. • A study to further refine loss and dilution assumptions in the Mineral Reserve model should be undertaken. SRK recommends the use of a Regularised Reserve Block Model in any future optimizations or feasibility studies. • Increasing the mining rate may improve the NPV. A better understanding of the potential for ramp congestion by increasing the size of the truck fleet, the availability of additional dig faces in the pit, along with capital implications should be developed.

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2 Introduction SRK Consulting (Australasia) Pty Ltd (SRK) was retained by the Amarillo Gold Corporation (Amarillo) to revise the 2017 Pre-Feasibility Study (PFS) update for the Posse Mine Project (PMP), located in the municipality of Mara Rosa in the state of Goiás, Brazil, 360 km to the north of the state capital, Goiânia.

Once constructed, the Posse Mine Project (the Project) will consist of an open pit gold mine and related processing facilities for approximately 23.8 million tonnes of Mill feed (dry base) at a rate of 2.5-3.0 million tonnes per year (Mtpa) depending on grind size utilized.

The PFS for the Posse Mine Project was prepared by Coffey Consultoria e Serviços Ltda (Coffey) in 2011. An update of the PFS was undertaken in 2017 by SRK Consultores do Brasil Ltda after completion of the geotechnical study for the definition of the final pit slope angles and the updating of the Mineral Resource estimate (MRE), which was concluded in January 2017 by Australian Exploration Field Services (AEFS), based on incorporation of new borehole data.

Since this update, Amarillo has engaged Whittle Consulting (Whittle) to strategically optimise the Project. SRK has applied the results of Whittle’s work to design a practical open pit and re-estimate the Mineral Reserve and update the project economics.

Amarillo has previously filed the following NI 43-101 technical reports which include Mineral Resource Estimates for the Project as follows:

• Caracle Creek International Consulting, 2008. Independent Technical Report and Preliminary Economic Assessment, Mara Rosa Gold Property, Goiás State, Brazil. Report prepared for Amarillo Gold Corporation dated 29 February 2008. • Hoogvliet Contract Services and Australian Exploration Field Services Pty Ltd, 2010. Independent Mineral Resource Estimate and Preliminary Economic Assessment, Posse Deposit, Mara Rosa, Goiás State, Brazil. Report prepared for Amarillo Gold Corporation dated 30 June 2010. • Hoogvliet Contract Services and Australian Exploration Field Services Pty Ltd, 2011. Report on Independent Site Visit and Resource Estimate. Posse Deposit, Mara Rosa, Goiás State, Brazil. Report prepared for Amarillo Gold Corporation dated 30 July 2011. • Coffey Mining Pre-Feasibility Study, Mara Rosa Project, Goiás State, Brazil. Report prepared for Amarillo Gold Corporation, dated 28 October 2011. • Australian Exploration Field Services Pty Ltd, 2016. Posse Deposit, Mara Rosa, Goiás State Brazil, Mineral Resource Update prepared for Amarillo Gold Corporation dated 21 July 2016. • SRK Consultores do Brasil Ltda, 2017 Updated PFS, Posse Mine Project – Mara Rosa GO prepared for Amarillo Gold Corporation dated 11 April 2017.

This report updates the previous 2017 PFS update.

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3 Reliance on Other Experts This PFS is a report in compliance with the methodology and format outlined in National Instrument 43-101, companion policy NI 43-101CP and Form 43-101 F1.

The Qualified Person for this report and for the disclosure of ore reserves is Anthony Stepcich of SRK.

Anthony Stepcich has relied on the following experts in the writing of this report:

• Keith Whitehouse of Australian Exploration Field Services of Bendigo, Australia, for the Mineral Resource statement • Frank Baker from Amarillo for the study process, processing plant and infrastructure components • Rich Peevers and Jason Pan from Whittle Consulting Pty Ltd for pit optimization and scheduling components.

The Mineral Resource estimate was prepared by Mr Keith Whitehouse, a competent and professional individual from Australian Exploration Field Services Pty Ltd, on behalf of Amarillo and is directed solely for the development and presentation of data with recommendations to allow Amarillo and current or potential partners to reach informed decisions. The information, conclusions and recommendations contained herein are based on work carried out by the author. The author has relied on his review of the site and site procedures in 2012 together with digital and hard copy data and information supplied by Amarillo, various published geological reports and discussions with representatives from Amarillo who are familiar with the Posse deposit and the area in general. The author has investigated the reports and other data listed in the ‘References’ section of this report pertaining to the Mineral Resource and is of the opinion that they are substantially accurate and complete.

In undertaking this update of the previous 2017 PFS report by SRK Consultores do Brasil Ltda, Anthony Stepcich has not yet undertaken a site visit as QP. The site visit was deferred and it is currently planned for Mr Stepcich to visit the Project in September/ October 2018. After the proposed site visit, this Report will be updated with any material changes and re-filed with SEDAR.

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4 Property Description and Location The Mara Rosa Property (also generally known and referred to as the Posse deposit) is located in Goiás state, central Brazil, approximately 6 km north of the town of Mara Rosa. The Posse deposit is centred at approximate Latitude 13° 58.395′ S, Longitude 49° 10.690′ W (Datum WGS84) or 696,900 mE, 8454,500 mN (Datum WGS84, Projection UTM, Zone 22 South), as shown in Figure 4-1. The Project encompasses a land area of 2,552 ha across three mining concessions plus numerous exploration leases in areas surrounding the Posse mine area.

Western Mining Corporation (WMC) operated a small open pit mine at the project site during the 1990s. Two pits, Posse South and Posse North, were developed over a five-year period. The ore, along with feed from the nearby Zacarias mine, was processed on site. The processing, beginning with heap leach and later Carbon-in-Leach (CIL), was conducted on approximately 10 ha of freehold property adjacent to the mining leases. Local infrastructure included adequate power and water to run a 600 tonnes per day CIL plant and heap leach operation.

As of November 2006, the mine and mill site had been reclaimed and no site infrastructure remained. According to Amarillo, the required remediation for mine closure had been met and accepted by the relevant government agencies. No significant environmental liabilities are known to exist at the former mine site.

WMC maintained a core logging and storage facility, sample preparation laboratory, assay laboratory, and office complex immediately north of the town of Mara Rosa. The facilities, which occupy 8 ha of freehold land, have been utilized by Amarillo during their exploration programs. As of July 2012, when Mr Whitehouse visited the project, the structures remain in excellent condition. The offices were utilized by Mr Whitehouse during his site visit and Amarillo staff during the drilling program, which finished in December 2012. Amarillo also owns two houses on contiguous pieces of land on São Paolo Street in the town of Mara Rosa.

Table 4-1 shows a list of current concessions and tenements owned by Amarillo that make up the Posse project. Amarillo has stated to SRK that all Amarillo concessions and tenements listed below are currently valid.

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Table 4-1: Concession and tenement schedule

Registered Area DNPM Number Status Tenement in: Partial Report in Remarks Project Owner (in ha) 1 861.241/1980 M. Concession Mine Suspension - Report Amarillo Mineração do Brasil Ltda. Mara Rosa 566.6 Submitted in 2017 - Awating 2 860.952/1980 M. Concession Amarillo Mineração do Brasil Ltda. Mara Rosa 1,000.0 Decision of DNPM - Requested 3 years 3 862.000/1984 M. Concession Amarillo Mineração do Brasil Ltda. Mara Rosa 986.0 Extension 4 862.021/2011 Tenement Sept, 14, 2015 July 17, 2018 RPP Submitted on July 16th Amarillo Mineração do Brasil Ltda. South Zacarias 768.1 5 862.714/2011 Tenement Sept, 24, 2015 July, 26th, 2018 RPP Submitted on July 26th Amarillo Mineração do Brasil Ltda. C. Verdes 1,762.7 6 862.715/2011 Tenement Sept, 24, 2015 July, 26th, 2018 RPP Submitted on July 26th Amarillo Mineração do Brasil Ltda. C. Verdes 780.5 7 862.719/2011 Tenement Sept, 24, 2015 July, 26th, 2018 RPP Submitted on July 26th Amarillo Mineração do Brasil Ltda. C. Verdes 1,987.8 8 862.720/2011 Tenement Sept, 24, 2015 July, 26th, 2018 RPP Submitted on July 26th Amarillo Mineração do Brasil Ltda. C. Verdes 1,970.1 9 862.721/2011 Tenement Sept, 24, 2015 July, 26th, 2018 RPP Submitted on July 26th Amarillo Mineração do Brasil Ltda. C. Verdes 1,955.7 10 862.722/2011 Tenement Sept, 24, 2015 July, 26th, 2018 RPP Submitted on July 26th Amarillo Mineração do Brasil Ltda. C. Verdes 1,719.0 11 861.028/2012 Tenement Feb, 24, 2016 Dec. 24, 2018 Amarillo Mineração do Brasil Ltda. C. Verdes 510.8 12 861.029/2012 Tenement Feb, 24, 2016 Dec. 24, 2018 Amarillo Mineração do Brasil Ltda. Mara Rosa 1,999.8 13 861.945/2013 Tenement Sept, 03, 2015 July. 03, 2017 Submitted in June 29, 2017 Amarillo Mineração do Brasil Ltda. Amaralina 1,645.0 14 861.946/2013 Tenement Sept, 03, 2015 July. 03, 2017 Submitted in June 29, 2017 Amarillo Mineração do Brasil Ltda. Mara Rosa 1,983.8 15 861.947/2013 Tenement Sept, 03, 2015 July 05, 2018 Submitted in July 05th, 2018 Amarillo Mineração do Brasil Ltda. Alto Horizonte 1,669.6 16 860.718/2013 Tenement May 02, 2016 March 02, 2019 Amarillo Mineração do Brasil Ltda. Amaralina 1,999.8 17 860.719/2013 Tenement May 02, 2016 March 02, 2019 Amarillo Mineração do Brasil Ltda. Amaralina 1,982.3 18 860.720/2013 Tenement May 02, 2016 March 02, 2019 Amarillo Mineração do Brasil Ltda. Amaralina 1,999.9 19 860.721/2013 Tenement May 02, 2016 March 02, 2019 Amarillo Mineração do Brasil Ltda. Amaralina 1,999.9 20 860.722/2013 Tenement May 02, 2016 March 02, 2019 Amarillo Mineração do Brasil Ltda. Amaralina 1,999.9 21 861.948/2013 Tenement Sept. 03, 2015 July. 05, 2018 Submitted in July 05th, 2018 Amarillo Mineração do Brasil Ltda. C. Verdes 1,954.4 22 860.514/2014 Tenement Aug 28, 2015 Jun 29, 2017 Submitted in June 29, 2017 Amarillo Mineração do Brasil Ltda. Mara Rosa 1,822.8 23 860.864/2016 Tenement Feb, 17, 2017 Dec 17th, 2019 Amarillo Mineração do Brasil Ltda. C. Verdes 1,971.8 24 860.865/2016 Tenement Feb, 17, 2017 Dec 17th, 2019 Amarillo Mineração do Brasil Ltda. C. Verdes 1,968.5 25 860.866/2016 Tenement Mar, 20, 2017 Mar, 20, 2020 Amarillo Mineração do Brasil Ltda. C. Verdes 1,987.5 26 860.867/2016 Tenement Feb, 17, 2017 Dec 17th, 2019 Amarillo Mineração do Brasil Ltda. C. Verdes 1,801.6 27 860.868/2016 Tenement Feb, 17, 2017 Dec 17th, 2019 Amarillo Mineração do Brasil Ltda. C. Verdes 1,997.0 28 860.869/2016 Tenement Mar, 20, 2017 Jan, 20, 2020 Amarillo Mineração do Brasil Ltda. C. Verdes 1,723.6 29 860.870/2016 Tenement Feb, 17, 2017 Dec 17th, 2019 Amarillo Mineração do Brasil Ltda. C. Verdes 2,000.0 30 860.871/2016 Tenement Feb, 17, 2017 Dec 17th, 2019 Amarillo Mineração do Brasil Ltda. C. Verdes 2,000.0

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Registered Area DNPM Number Status Tenement in: Partial Report in Remarks Project Owner (in ha) 31 860.100/2017 Tenement Mar 02, 2017 Jan, 02, 2020 Amarillo Mineração do Brasil Ltda. Mara Rosa 878.1 32 860.101/2017 Tenement April 06, 2017 Feb 06, 2020 Amarillo Mineração do Brasil Ltda. Mara Rosa 1,880.0 33 860.102/2017 Tenement April 06, 2017 Feb 06, 2020 Amarillo Mineração do Brasil Ltda. Mara Rosa 1,853.4 34 860.103/2017 Tenement Mar. 02, 2017 Jan, 02, 2020 Amarillo Mineração do Brasil Ltda. Mara Rosa 1,074.5 35 860.104/2017 Tenement May 22, 2017 March 22, 2020 Amarillo Mineração do Brasil Ltda. Mara Rosa 874.6 36 860.105/2017 Tenement May 22, 2017 March 22, 2020 Amarillo Mineração do Brasil Ltda. Mara Rosa 1,632.5 37 860.106/2017 Tenement Mar, 02, 2017 March 02, 2020 Amarillo Mineração do Brasil Ltda. C. Verdes 839.0 38 860.107/2017 Tenement Mar, 02, 2017 March 02, 2020 Amarillo Mineração do Brasil Ltda. Amaralina 1,632.3

Total in ha. 61,178.5

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Figure 4-1: Location of Amarillo’s Mara Rosa properties

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5 Accessibility, Climate, Local Resources, Infrastructure and Physiography Details of the of surface rights, availability of power, water, labour and both waste disposal and process plant locations are discussed in more detail in separate relevant sections of this report. Suffice it to note that the Project is not seriously constrained by space or other factors necessary for mining activities.

5.1 Accessibility The Municipality of Mara Rosa is located 356 km north of Goiânia in the Porangatu Microregion, 11 km west of the Belém-Brasília highway, between the basins of the Araguaia River and the Tocantins River (Figure 5-1). According to a recent estimate, Mara Rosa has a population of approximately 12,000, of whom 10,000 reside in the town.

5.2 Climate Average annual rainfall is approximately 1,500 mm, resulting in a relatively wet climate. The year is defined by two principal seasons, a dry season from April to September and a wet season from October to March. The mean temperature is 24°C during the dry season and 28°C during the wet season. Annual temperatures typically range from approximately 4°C to 45°C. The climate does not impose any limitations on exploration or potential mining operations and they can continue throughout the year.

5.3 Local Resources Local facilities include several public and private elementary and high schools, two hospitals, a public health centre, three banks, several small hotels, three petrol stations and numerous shops. Agriculture (corn, rice, manioc, sugarcane, soybeans, and bananas) and cattle ranching are the primary commercial activities in the region. Mara Rosa is a regional support community for these activities.

5.4 Infrastructure The municipality has an excellent network of local farm roads, the majority of which are unpaved but generally in good condition. The municipality is also serviced by an 800 m long, unpaved airstrip. Access to Mara Rosa is via Federal Highway BR-153, the main north–south highway in central Brazil leading to the city of Belém at the mouth of the Tocantins River. Mara Rosa is 356 km, or 4 hours driving time, north from the state capital of Goiânia, and 320 km, or 4 hours driving time from the national capital, Brasilia. Highway communications with Goiânia are made by GO-080/ Nerópolis/ São Francisco de Goiás/ BR-153/ Jaraguá/ GO-080/ Goianésia/ Barro Alto/ GO-342/ BR-153/ Uruaçu/ Campinorte/ GO-239 (Figure 5-1).

Electric power is supplied by CELG, the State of Goiás Energy Authority. The local electricity grid has an installed capacity of 14 MW supplied to the area via a 64 kV line. Should the proposed mine be developed, a new transmission at 138 kV will be installed to supply the mine as stated in the 2011 PFS. The water supply is metered and is provided by SANEAGO, the state water company. Water for the Posse mine site as well as ranches in the surrounding region is derived from a combination of local streams and artesian wells. Telephone service, both local and international, is provided by TELEGOIAS. Cellular telephone service is available in the area.

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5.5 Physiography The region is characterised by tropical savannah of low to moderate topographic relief ranging from approximately 400 m to 500 m above sea level (ASL). The town itself has a mean elevation of 520 m ASL. Much of the area has been cleared for farming and as a result is open savannah grassland. Trees occur along the abundant water courses.

Figure 5-1: Mara Rosa and surrounding towns

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6 History This section summarizes the work carried out prior to the release of this report.

Evidence of small-scale surficial-alluvial mining along the Rio do Ouro in the historical Amaro Leite area indicates mining activity in the Mara Rosa district dates to the mid-1700s. More recent activity dating from the early 1970s to early 1980s began with the successful discoveries by INCO (now Vale S.A. or Vale) of the Chapada gold-copper and Crixás gold deposits. These deposits are located approximately 30 km and 100 km to the southwest of the town of Mara Rosa, respectively.

During the early 1980s, BHP-Utah Mines (now BHP Billiton Limited), through its subsidiary Mineração Colorado Ltda., initiated a grassroots reconnaissance program that covered the Chapada district and the Mara Rosa area, and eventually led to the discovery of the Posse gold and Zacarias gold-silver- barite deposits. From 1981 to 1987, BHP completed 12,300 m of diamond and reverse circulation (RC) drilling at Posse and Zacarias. At Posse, a 107 m exploration shaft was sunk and 400 m of lateral drifting was completed to test mineralization.

As a result of Brazilian restrictions on foreign ownership in 1988, BHP chose to joint venture the Mara Rosa properties with Western Mining Corp. In 1990, WMC set up a subsidiary, Mineração Jenipapo S.A. (MJSA), to acquire a 100% interest in Posse, and to explore, develop, and operate the asset. The Posse mine was opened in 1992 and operated until July 1995, during which time the Posse North and Posse South pits were developed. The on-site mill processed approximately 750,000 tonnes of ore grading a combined 3.5 g/t Au. Zacarias, which was significantly higher grade, operated at roughly the same time as Posse and was processed through the same mill.

In order to provide cash flow for its activities in Brazil, WMC focused much of its attention on development of the Posse and Zacarias mines between 1990 and 1995. This work is understood to have been completed as a result of a corporate decision to make each business unit self-funding and to encourage efforts to develop known deposits. In addition, efforts to replace mined reserves were directed toward both the Eastern and Central Belt exploration targets generated previously by BHP as well as new targets identified to the east and north of Mara Rosa.

By June 1995, a combination of factors, including low gold prices, the exhaustion of reserves at the higher grade Zacarias deposit, and the failure to discover any additional, near-surface reserves, caused WMC to discontinue mining and exploration activities at Mara Rosa. As the primary exploration objective had been the discovery of near-surface mineralization that could be fast-tracked into production, most of the exploration targets identified by BHP and WMC had only been evaluated to depths within approximately 50 m from the surface.

Upon suspension of its mining and exploration activities, WMC was approached by several companies interested in exploring the property under lease-option agreements. The Zacarias deposit and the rights to its tailings were eventually sold to Minere Mineração Ltda (Minere), a small Brazilian company interested in exploiting the deposit’s very high barite content. The project has since been on-sold to a company called Baribras Mineração Ltda.

In 1996, Barrick do Brasil (Barrick) completed a full due diligence study of the remaining Posse project concessions (the Eastern Belt claims). The due diligence involved a team of at least 14 people and a significant program of test sampling, re-logging of core, soil sampling, reinterpretation of geophysics, and an estimate of the mineral resource for the Posse deposit. Although this program subsequently led to a preliminary offer by Barrick to purchase the property in full, negotiations stalled prior to execution of the agreement. Barrick provided WMC with a copy of its due diligence report and related correspondence after the failure to execute a deal.

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Following Barrick’s withdrawal, Metallica entered into negotiations with WMC for the purchase of the Eastern Belt properties, and in November 1997, successfully completed an agreement, which called for a total purchase price of US$1.5 million. As part of the previous buy-out agreement between BHP and WMC, BHP held a 1% net smelter return (NSR) royalty interest on the property. This now sits with Royal Gold after a royalty portfolio sale by BHP. Euro-Nevada Gold Corporation (later absorbed into Newmont) held an additional 1% NSR royalty. This now sits with Franco Nevada Corporation, after this royalty-focused corporation folded out of Newmont.

Following a compilation of data and a review of the project, Metallica completed a systematic soil geochemistry and geological mapping program northeast of the Posse deposit. Induced polarisation (IP) and ground magnetic geophysical surveys were completed over some of the more promising areas. Metallica suspended exploration operations in September 1998 and placed the Project on care and maintenance. In 2001, Metallica revisited the Project and completed a review of the regional potential. At this time, five holes, totalling 940 m, were drilled into three separate targets on the northern extensions to the Posse mine trend. Following this work, a corporate decision was made to focus on properties in Mexico and Chile and Metallica decided to sell the Project.

Amarillo Gold Corporation (Amarillo) visited the Project in August 2003 and in October 2003 signed a letter of intent with Metallica to purchase MBL and 100% of the Mara Rosa project. The Project remains subject to the 1.0% NSR royalty to Franco Nevada Corporation and a further 1.0% royalty to Royal Gold. During the time that Amarillo has controlled the property, the Company has undertaken considerable work to define the extent and nature of the Posse deposit with the aim of developing the primary or fresh (non-oxidised) mineralization.

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7 Geological Setting and Mineralization The Mara Rosa District is situated within the Goiás Magmatic Arc (GMA) which forms part of the Tocantins physiographic province, an intercratonic mobile belt that separates the Amazonas and São Francisco cratons, located to the northwest and southeast respectively. The GMA is a 100 km wide, northeast trending granite-greenstone terrane that extends for approximately 700 km. The geology in the Mara Rosa District is principally delineated by three northeast striking, moderately to steeply northwest dipping belts of metamorphosed volcanosedimentary and associated intrusive rocks. These belts, referred to as the Western, Central, and Eastern Belts, are separated by broad zones of tonalitic orthogneiss.

The Eastern Belt is bounded to the southeast by the Rio dos Bois fault, which also defines the south eastern limit of the GMA.

Amarillo’s land position within the Mara Rosa District primarily covers the Eastern Belt greenstone assemblage with some coverage of the Western and Central belts as well. The Eastern Belt, has a maximum thickness of 6 km, generally strikes to the northeast and dips moderately to steeply to the northwest. Surface topography is characterised by moderate relief and locally dissected drainages that follow lithologic or structural weaknesses. The depth to fresh bedrock is generally shallow, ranging from 0 m to 15 m. The upper portion of the weathered profile consists of clay-rich latosol and saprolite derived from the underlying bedrock.

Rocks of the Eastern Belt are locally intruded by quartz-feldspar-muscovite and biotite granitic rocks and associated aplite and pegmatite dykes, small stocks and dykes of hornblende, biotite and magnetite diorite, and, in its north-central portion, a large body of hornblende-plagioclase gabbro. All units exhibit varying degrees of foliation that typically range from weak to moderate, and generally intensify along sheared contacts. The tonalitic orthogneiss which separates the Eastern and Central Belts is composed of coarse-grained plagioclase, hornblende, and biotite with localised patches of biotite schist near its contact with the Eastern Belt.

Structurally the Eastern Belt is dominated by well-developed, penetrative foliation that strikes 30°–50° and dips 40°–70° northwest – an orientation subparallel to stratigraphy. Major structural systems include 50°–65° striking shears and thrusts and associated drag folds. Shears are most commonly developed along zones of elastic disparity such as lithologic contacts. Shear sense is typically reverse-dextral oblique although a sinistral sense is locally observed. A second set of structures consist of late cross cutting northwest to east–northeast striking brittle faults and fractures that locally offset stratigraphy in apparent dextral strike-slip sense.

Uranium-lead isotopic age determination of zircons from some of the principal lithologic units within the district indicates timing of initial rock formation for both the belt rocks and the tonalite gneiss to be between approximately 870 Ma to 850 Ma (Viana, 1995). Subsequent amphibolite facies metamorphism is estimated to have occurred between 700 Ma and 600 Ma based on U-Pb and Rb-Sr dating of recrystallised titanate. The latter date corresponds to peak metamorphism related to the Brasiliano orogenic event.

Several significant mineral deposits occur within 50 km of Mara Rosa town including the Posse gold deposit, the Zacarias gold-silver deposit and the Chapada copper-gold deposit, together with numerous historical prospects and small-scale historical mines locally known as garimpos.

7.1 Local Geology The Posse deposit occurs in a regional thrust that probably acted as one of the primary dewatering conduits during the Neo-Proterozoic Brasiliano orogeny. The geophysical, geological and geochemical data available demonstrate that the Posse deposit occurs within a 50 km long shear zone

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with potassium alteration and lower order gold-copper-molybdenum mineralization. The Posse deposit has grey gneiss in the hanging wall of the fault and amphibolite, ‘greenstone’ in the footwall. Shearing of the Grey Gneiss has resulted in the formation of a distinct lithologic unit, a quartz-feldspar-mica schist (Posse Schist) that is characteristic of the Posse ore zone. This unit has been identified in several other areas including the Posse footwall and on strike extensions of the Posse Ore Zone to the northeast. Shearing is most intense in the footwall. It is speculated that the rheological contrast between the hanging wall and footwall rock types captured the regional thrust (movement west to east) for a 2 km segment of the shear. It is also possible that the chemical contrast between the hanging wall and footwall rocks may have aided in focusing mineralizing fluids. Observations from drill core suggest that an earlier potassic event with quartz veining, chalcopyrite, molybdenum, biotite and K-feldspar was followed by a later phyllic (sericite) event with pyrite, iron- telluride, and gold. Gold occurs as native gold and also with telluride and pyrite.

In general, mineralization at Posse is developed along a 050°–065° striking fault zone. Mineralization tends to be strongest within mylonitic zones that follow more northerly striking (approximately 030°-050°) shear strands and dilatant jogs that obliquely transect the contact between the hanging wall and footwall rocks.

The mineralization envelope at Posse is about 30 m thick and over 1 km long (Figure 7-1). It has a mylonitic appearance which is most noticeable in the footwall where shearing is the most intense. Higher intensity of shearing is associated with increased sulphide mineralization (up to about 4%), and a slight increase in metamorphic grade from greenschist to high greenschist facies in the hanging wall through to high greenschist/ low amphibolite facies in the footwall (biotite flakes and garnet alteration). Higher gold values are associated with increasing intensity of shearing and higher levels of sulphide mineralization.

Aside from the slight increase in metamorphic grade, there appears, based on inductively coupled plasma (ICP) analyses obtained from the 2005/ 2006 drilling program, to be a chemical difference in lithology between the hanging wall and footwall; however, this is not visually obvious.

The shear zone may be more complicated than a simple main shear near the footwall with gradually decreasing intensity towards the hanging wall. Based on geochemical evidence there is some reason to believe that different portions of the shear zone were active at different times.

A thin basaltic dyke which does not offset the mineralization has been intersected in some drill holes.

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Figure 7-1: Geology of the Posse deposit

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8 Deposit Types Several significant mineral deposits occur in the Mara Rosa region including the Posse gold deposit, the Zacarias gold-silver-barite deposit and the Chapada copper-gold deposit, in addition to numerous historical prospects and garimpos. These are listed in Table 8-1.

Table 8-1: Significant deposits in the Mara Rosa region

Deposit Deposit Class References Posse Au Metallica data (Mara Rosa files) and Shear-hosted mesothermal lode-gold. (Eastern Belt) Amarillo website Zacarias Au-Ag-Ba Stratiform syngenetic exhalative or shear WMC data (MR files); Poll, 1994. (Central Belt) related epigenetic high sulphidation? R Shaw/ M Petersen Chapada Cu-Au Volcanogenic exhalative? Wall rock Kuyumjian, 1991; Richardson, et al., (Eastern Belt) porphyry copper system? 1986; 1988

The Posse deposit is hosted in a regional thrust that probably acted as one of the primary dewatering conduits during the Neo-Proterozoic Brasiliano orogeny. The geophysical, geological and geochemical data available demonstrate that the Posse deposit occurs within a 50 km long structural zone with potassium alteration and lower order gold-copper-molybdenum mineralization. The Posse deposit has a hanging wall of grey gneiss and the footwall of amphibolites, ‘greenstone’, and it is speculated that the rheological contrast between the two rock types captured the regional thrust (movement west to east) for a 2 km segment. It is also possible that the chemical contrast between the acidic hanging wall and basic footwall may have aided in focusing the mineralizing fluids. Observations in the core suggest that an earlier potassic event with chalcopyrite, molybdenum, quartz veining, biotite and K-feldspar was followed by a later auriferous phyllic event with pyrite occurs free as well as associated with the telluride and pyrite.

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9 Exploration During the 1990s WMC operated a small open pit mine at the Project. The Posse South and Posse North pits were developed over a 5-year period and oxide ore was processed on site. The mine and mill site were reclaimed and no site infrastructure remained by November 2006. No significant environmental liabilities are known to exist at the former mine site and it is understood that the required remediation for mine closure had been met and accepted by the appropriate government agencies. Numerous drilling campaigns have been completed on the property:

• BHP Billiton: 1982–1987 • WMC: 1988–1995 • Amarillo: 2005–2006 • Amarillo: 2008 • Amarillo: 2010–2011 • Amarillo: 2011–2012.

In all, the drill hole data base contains 346 drill holes totalling 44,739 m of drilling.

During the period from late 2012 until June 2018, no drilling was carried out or samples submitted for assay. Amarillo started a 56-hole drilling program at the Project in June 2018. The program consists of 30 reverse circulation (RC) drill holes, 18PRC001 - 18PRC030, for a total planned length of 3,150 m and 26 diamond drill holes, 18P047–18P072, with a total planned length of 7,345 m. At the time this report was compiled 14 of the RC holes and 15 of the diamond holes had been completed. Work is currently underway on four of the diamond holes. The drill program is expected to be completed towards the end of 2018. A plan showing the location of the current and planned drilling is shown in Figure 9-1.

No results of the recent drilling have been provided and this report is based on data gathered prior to May 2018. All exploration prior to June 2018 is covered in the 2016 Resource report and earlier reports referred to in Section 2.

Figure 9-1: Location of drill holes in the 2018 drill program

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10 Drilling New resource drilling work was commenced at the Project in June 2018. No results from this work have been received. There was no drilling conducted from late 2012 until June 2018 and the drilling work on which this report is based was discussed in the 2016 Report where the drilling process and history has been detailed.

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11 Sample Preparation, Analyses, and Security The sample preparation, analyses and security has not changed since the 2016 Report where the procedure is described in detail.

The initial sample preparation is carried out in Goiania by ACME Laboratories (ACME). The procedure at ACME includes the following steps:

• Sorting and checking against the requisition sheet • Drying at 60°C • Washing with a granite wash to scour the equipment before the first sample is crushed • Crushing of the samples to 80% passing 10 mesh (2 mm) • Samples homogenised and riffle split to 250 g subsample • Subsamples are pulverised to 85% passing 200 mesh (75 micron) • Equipment cleaning by brush and pressurised air • A granite wash is used to scour the equipment after high grade samples, between changes in rock colour, and at the end of each file.

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12 Data Verification Extensive data validation was carried out in the lead up to the estimation of the 2016 Mineral Resource update; this is detailed in the 2016 Resource report.

Following the release of the 2016 Mineral Resource update, further work was carried out to allow better control of the specific gravity (SG) used to calculate tonnage from volume in the model. In the 2016 Mineral Resource update, based on a review of geophysical data, an SG of 2,730 kg/m3 was applied to the whole orebody. In the PFS completed in 2011, SG values of 1,800 kg/m3 for soil and 2,400 kg/m3 for weathered rock were also used, but the basis for subdividing the deposit into soil, weathered and fresh rock was, in places, inconsistent. This was a function of historical lithology coding.

Discussions with Amarillo staff led to a concerted effort to recode the lithology data to allow the different SG zones to be properly modelled. Updated wireframes showing base of soil and base of weathering were examined on a section by section basis to ensure that fresh rock did not overlap weathered or soil and that weathered rock did not overlap soil. This information was then fed into the block model, which resulted in a slight adjustment in tonnage of the orebody. Apart from alteration to the SG values used in the Mineral Resource estimate and to the cut-off grade at which the Mineral Resource is reported, all input data and modelling parameters are the same as those used in the 2016 Resource report.

The update to the Mineral Resource, which is included as part of this report, has occurred as the result of a re-examination of the topographic surface used in the 2016 and 2017 Mineral Resource reports. The topographic surface used for the earlier Mineral Resource reporting, updated in 2015, with the inclusion of GPS topographic survey data, within the footprint of the previously mined area represents a backfill surface. Historical survey records have now been used by Amarillo to determine the maximum extent of mining and a new surface which incorporates this information has been generated. The volume between the backfilled surface and the historical final mined surface has been recognized as backfill material. The surfaces used for this work were provided by Amarillo in the form of DXF files. These were imported as wireframes into the modelling software (Micromine, 2018 SP2).

Section by section comparisons between the topographic surface used for the 2016 and 2017 resource reports, the backfilled surface and the historical final mined surface have been made. There is little or no difference in the topographic profile between these surfaces outside the footprint of the Posse pits. Within the footprint the backfill surface was a good match to the previously used topographic surface. Similarly, the historical final mined surface was generally, as expected, below the backfilled surface. There were zones at the edge of the pit footprint where the mined surface was higher than the backfilled surface. This material, approximately 69,000 m3, is assumed to have been used for some of the backfill material. However, there is approximately 395,000 m3 of backfill material in the historical pit and the source of 326,000 m3 of the backfill material has not been determined.

It has been speculated that historical waste dumps may have been used. However, this has not been verified as these areas have not been well surveyed due to dense vegetation. Further work is planned to define the source of this material; however, because the nature of the backfill is uncertain, all blocks in the block model within this volume have been excluded from the current Mineral Resource.

As there had been a number of updates to the software used for Mineral Resource modelling and reporting since the Mineral Resource was modelled in 2016, a number of modelling runs were conducted to confirm that the values recorded in the model were the same as those that would be estimated with the latest versions of the software. All previous reported results were duplicated using the latest versions of the software.

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13 Mineral Processing and Metallurgical Testing The authors are not aware of any metallurgical study work, which has been undertaken since the report, 2017 Updated Pre-Feasibility Study Posse Mine project – Mara Rosa GO, which has been filed on SEDAR by Amarillo on 4 May 2017.

The proposed process is summarized as follows and presented in Figure 13-1.

The processing plant consists of a conventional crushing circuit (including tertiary crushing) followed by primary and then secondary milling in closed circuit with cyclones. The final pulp at a P80 of 45 µm is pre-oxidised in agitation tanks using oxygen gas from an oxygen plant at a high pH of 12 for a total of 12 hours to oxidise tellurides to enable successful cyanidation of the gold. The pulp is contacted with cyanide and activated carbon in a typical CIL circuit of six agitated tanks for a total of 24 hours.

Loaded carbon is extracted daily from the CIL circuit and processed in a typical Zadra-style elution circuit at up to 140°C with a 4 tonne capacity. The eluted solution is passed continuously to the electrowinning cells until efficient desorption has been achieved. At intervals, the gold is removed from the cells and smelted into doré bars for refining and sale. The activated carbon is regenerated in a gas fired rotating kiln before being sent back to the CIL circuit.

The tails from the CIL circuit is thickened to recover some of the solution before the thickened pulp is subjected to detoxification with SO2/ air and a copper sulphate catalyst to destroy free cyanide before being pumped to a tailings storage facility (TSF). The supernatant from TSF is recirculated to the plant under conditions of zero discharge.

Figure 13-1: Proposed flowchart for the operation

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14 Mineral Resource Estimates The Mineral Resource estimate for Posse has been updated by Gregory Keith Whitehouse, BSc (Geology), MAusIMM(CP) of Australian Exploration Field Services Pty Ltd (AEFS), to reflect additional information on the extent of historical mining and the backfilling of the Posse pits. This information was not included in the Mineral Resource model and estimate reported in 2016 or the update to the 2016 estimate reported in 2017. In addition, recent pit and mine optimization work on the Project has indicated that some material will be economic to mine at a cut-off grade of 0.216 g/t Au. Accordingly, the cut-off grade used for the Mineral Resource estimate has been lowered to 0.2 g/t Au to ensure that all blocks which may be bought into Mineral Reserves are included in the Mineral Resource.

The 2018 Mineral Resource estimate is summarized in Table 14-1.

Table 14-1: Mineral Resource estimate - 2018

Grade Ounces Category Tonnes (Mt) (g/t Au) (Moz) Measured Mineral Resource 12 1.4 0.56 Indicated Mineral Resource 19 1.2 0.71 Total - Measured and Indicated Mineral Resource 31 1.3 1.3 Inferred Mineral Resource 11 0.92 0.33 Source: AEFS, 2018. Note that tonnes, grade and ounces in the 2018 Mineral Resource estimate shown in Table 14-1 have been reported to two significant figures only to reflect the uncertainty inherent in any Mineral Resource estimate. A cut-off grade of 0.2 g/t Au has been used for the Mineral Resource estimate.

The Posse deposit is situated near the town of Mara Rosa in Goiás State, Brazil. Australian Exploration Field Services Pty Ltd (AEFS) was retained by Amarillo to visit the site, review procedures, validate and where appropriate correct the drill hole database for the Posse deposit. Subsequently, the deposit was modelled by Mr Whitehouse and an estimate of the gold Mineral Resource was made in 2016. This work was reported in an NI 43- 101 compliant Technical Report dated July 2016.

The historical data used in the Mineral Resource estimate was detailed in the NI 43-101 report, dated 30 June 2010, titled Independent Mineral Resource Estimate and Preliminary Economic Assessment, authored by Hoogvliet Geological Services and AEFS (Hoogvliet & Whitehouse 2010). Drilling conducted by Amarillo between November 2010 and March 2011 was detailed in the NI 43-101 report, dated 30 July 2011, titled Report on Independent Site Visit and Resource Estimate, authored by Hoogvliet Geological Services and AEFS (Hoogvliet & Whitehouse 2011). Results from a further series of 59 diamond drill holes completed between June 2011 and December 2012 were discussed in Section 13 of the 2016 Mineral Resource estimate.

In July 2012, Mr Whitehouse undertook a site visit and review of site procedures which supports the 2016, 2017 and 2018 reports. There was no resource drilling from the end of 2012 until May of 2018. The resource definition drilling which was started in May 2018 has not been included in this report.

The 2016 estimate was updated by Mr Whitehouse in 2017 to reflect work carried out by Amarillo which improved understanding of the oxidation zoning of material in the model and an improved understanding of the economics of mining, which led to a reduced, lower cut-off grade for the Mineral Resource – from 0.5 g/t Au to 0.35 g/t Au. The updated report was included as part of the 2017 Updated Pre-Feasibility Study, Posse Mine Project – Mara Rosa/ GO by SRK Consultores do Brasil Ltda, dated April 2017.

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It was subsequently recognised that there was backfill material within the footprint of the historically mined pit which was inappropriately classified. As a result, a decision was made to update the Mineral Resource to exclude material identified as backfill. Surfaces compiled by Amarillo representing the backfilled profile and the historical final mined profile were used to identify the portions of the 2017 Mineral Resource affected by this reclassification.

At that same, time the cut-off for the Mineral Resources was dropped to 0.2 g/t Au to ensure all blocks likely to be brought into Mineral Reserve under the economic scenario adopted for this PFS update were within the Mineral Resource.

14.1 Historical Reports The Mineral Resource estimate in this report represents the seventh independent Mineral Resource estimate completed on behalf of Amarillo for the Posse deposit. The historical reports were:

1 Estimation of an Inferred Mineral Resource estimate by CCIC in March 2007. 2 An updated resource estimate which complied with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Mineral Reserves Definitions Guidelines by CCIC in February 2008. 3 HCS and AEFS completed Mineral Resource estimates in 2010 and 2011. The 2011 Mineral Resource estimate was used as the basis of the Pre-Feasibility Study conducted by Coffey Mining in 2011. 4 An update to the Mineral Resource which incorporated additional drilling and improved downhole surveys and surface topography was completed by AEFS in July 2016. 5 An update to the 2016 Mineral Resource which incorporated improved delineation of the oxidation zones within the resource and a reduced cut-off grade of 0.35 g/t Au, based on the results of pit optimization work was completed by AEFS in January 2017 and incorporated into an Update to the Pre-Feasibility Study by SRKBR in April 2017.

All resource reporting by HCS and AEFS has been in accordance with the CIM Mineral Resource and Reserve reporting standards current at the time the reports were written.

14.2 2018 Mineral Resource

14.2.1 Data Used Data used for modelling consisted of a set of drill collars, downhole surveys and assay data together with wireframe boundaries, which defined Hanging Wall, Main and Footwall zones of the deposit. The wireframes were developed by AEFS using the latest available set of drill hole data and are unchanged from the 2016 Mineral Resource estimate.

Additional wireframes representing the base of soil and the base of weathered rock were developed and used for the 2017 Mineral Resource estimate to assign appropriate density values to blocks in the weathered zone.

The surface topography used was the same as that used in the 2016 and 2017 Mineral Resource model; however, within the footprint of the historical pit, a backfilled and a historical final mined surface have both been modelled. The surfaces represent historical work completed by WMC. This information was provided by Amarillo and was validated by comparison with historical data held on file by AEFS.

A raw block model was generated in the WGS84, UTM Zone 22 S coordinate system with primary blocks of 25 m (E), 25 m (N) and 10 m (RL), sub-blocked to 5 m × 5 m × 5 m to fit the mineralized

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domain, topography and other wireframes. The blocks in this model were coded according to mineralized zone.

A default SG value of 2,730 kg/m3 was used for the SG of fresh rock. This is the same value as used in the 2010 and 2011 reports and verified by the downhole geophysics discussed in Section 13.1.1 of the 2016 Resource report. Wireframes representing the base of soil and the base of weathered rock were used to indicate which portion of the model should be assigned a density of 1,800 kg/m3 (Soil) and 2,400 kg/m3 (weathered rock). Portions of the model which fell between the backfilled surface and the historical final mined surface were assigned an SG of 2,000 kg/m3.

In the 2016 Mineral Resource model, the variography used in the 2010 and 2011 resource estimates was replaced with updated work, which generated separate semi-variograms for each of the three modelled domains, Hanging Wall, Main and Footwall. This same variography was used for both the 2017 and 2018 reports. Estimation of tonnage and grade of blocks within the constraining wireframes used a Multiple Indicator Kriging (MIK) algorithm. The data was then flagged to indicate confidence in the result and reported using a 0.2 g/t Au cut-off as Measured, Indicated and Inferred Mineral Resource blocks. All blocks within the volume constrained by the backfilled and historical final mined surface were excluded from the resource to reflect the uncertainty regarding the material in this zone. The categories ‘Measured, Indicated and Inferred’ have the same meaning as in CIM guidelines.

The steps taken to generate the final model are discussed in more detail below.

14.2.2 Wireframing Initial modelling work consisted of defining a set of consistent sections to be used in the development of wireframes based on drill hole intersections of geology and grade. The sections were oriented at an azimuth of 45° to the UTM grid so as to be at right angles to the strike of the mineralization.

As with earlier Mineral Resource estimates, three mineralization domains, equating to grade zones were recognized, Hanging Wall, Main and Footwall. The grade in both the Hanging Wall and Footwall was lower than the grade in the Main Zone and neither the Hanging Wall nor Footwall zones are continuous. Occasional splays of Main Zone into the Hanging Wall have been incorporated in the Hanging Wall domain as they are not well defined by drilling.

A late, unmineralized basaltic dyke, which had been encountered in a number of drill holes, was modelled and incorporated as a separate wireframe, which transects the Posse mineralization.

During construction of the wireframes the boundaries of the Hanging Wall and Footwall zones were deliberately allowed to intrude into the Main Zone. After the initial wireframes were built, they were clipped using the wireframe Boolean functions in Micromine to produce wireframes that had no under or overlap. The wireframes are shown in Figure 14-1, looking northeast; Figure 14-2, looking north and Figure 14-3, looking down.

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Figure 14-1: Posse wireframes and drill holes looking north east Note: The basalt dyke which transects the orebody is shown in light purple.

Figure 14-2: Posse wireframes and drill holes looking north along the strike of the basalt dyke

Figure 14-3: Posse wireframes and drill holes looking down

Additional wireframes representing the top of saprolite and top of fresh rock were constructed based on information provided by Amarillo. These surfaces generally sit one below the other and under the topography; however, in some areas they coalesce as there are areas with no soil and saprolite at surface, other areas have a deeper soil profile and a thinner saprolite profile. With the recognition of the historical final mined surface and the backfilled surface, additional wireframe surfaces were

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constructed to record this information and to allow the block model to be coded to indicate which zone blocks belong to. A long section showing the various surfaces is shown in Figure 14-4.

Figure 14-4: Surfaces used to control zonation in the block model

The volume of material affected by the volume between the backfilled and historically final mined surfaces is shown in yellow in Figure 14-5.

Figure 14-5: Volume, in yellow between backfill and mined surface in relation to the orebody

14.2.3 Compositing and Statistics Gold assay data was coded according to mineralised domain and was then composited to 1 m downhole intervals for modelling. This reduced the statistical variance of the grades and no further top-cutting of grade data was required. Key statistical parameters of the raw and composited data are shown in Table 14-2 and Table 14-3.

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Table 14-2: Summary statistics for raw assays in the mineralised zone

Raw Assays All Zones HW Main FW Normal Statistics Minimum 0.0025 0.0025 0.020 0.0025 Maximum 90.00 41.17 90.00 14.40 Number of points 12896 4815 6579 1502 Mean 1.02 0.47 1.58 0.33 Variance 5.15 2.33 7.65 0.42 Standard deviation 2.27 1.53 2.77 0.65 Median 0.45 0.21 0.87 0.0.18 Coefficient of variation 2.22 3.23 1.75 1.95 Outliers 1214 469 628 172 Sichel's T-Estimator 0.96 0.38 1.47 0.31

Table 14-3: Summary statistics for 1 m composite assays in the mineralised zone

1 m Composites All Zones HW Maine FW Normal Statistics Minimum 0.0025 0.0025 0.0025 0.0025 Maximum 41.17 41.17 41.12 13.52 Number of points 13073 4926 6306 1665 Mean 0.95 0.44 1.54 0.32 Variance 3.30 1.58 4.83 0.37 Standard deviation 1.82 1.26 2.20 0.61 Median 0.41 0.22 0.90 0.18 Coefficient of variation 1.91 2.85 1.43 1.91 Outliers 1219 484 583 183 Sichel's T-Estimator 0.96 0.37 1.46 0.29

A detailed breakdown of the statistics associated with each domain (HW, Main and FW) used in modelling was presented in both the 2016 and 2017 Mineral Resource reports. The drilling data is unchanged for this report as a result a summary only is presented.

Probability plots and histograms show that there are mixed populations of data in each of the Hanging Wall, Main and Footwall zones. Histograms showing the data distribution of the 1 m composites in each of the Hanging Wall, Main and Footwall zones are shown in Figure 14-6 to Figure 14-8.

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Figure 14-6: Grades of 1 m composites in the hanging Wall Zone

Figure 14-7: Grades of 1 m composites in the Main Zone

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Figure 14-8: Grades of 1 m composites in the Footwall Zone

14.2.4 Variography Semi-variograms of the median (50% cumulative value) were defined for each of the three zones (HW, Main & FW). From the variography, the anisotropy was defined as azimuth 228°, plunge 8° and dip -46.5° (left hand rule). This set of directions, which is an accordance with the known geological structure, was confirmed by using an MIP (Maximum Intensity Projection) of the 1 m composites in 3D space to define the directions in the model that showed the directions of highest grade continuity (E J Cowan 2014).

Note that the semi-variogram map in Figure 14-9 also shows that there is a direction of short range continuity at azimuth 247° with an anisotropy which was defined as azimuth 247°, plunge 21° and dip -44.5°. This fits with the semi-variogram parameters used in the 2011 PFS and is currently understood to be related to the high grade portion of the data (which forms 27% of the data points in the Main Zone) with a mean grade of 2 g/t Au, as shown in the probability plots.

Figure 14-9: Semi-variogram map of 1 m composites in the mineralization zone

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Downhole semi-variograms were used to define the nugget for each of the three zones, followed by the generation of 3D semi-variograms for each of the three components of the anisotropy for each of the zones as shown in Figure 14-10.

Figure 14-10: Directional semi-variograms for the three mineralization zones

The additional set of semi-variograms in Figure 14-11 was modelled to define the short range spatial relationships of the higher grade portion of the Main Zone (Main Hi). The high grade portion of the Main Zone comprises 27% of the data from which the probability plot decomposition had a mean grade of 2 g/t Au. From the semi-variogram, the anisotropy was defined as azimuth 247°, plunge 21° and dip -44.5° (left hand rule). These directions match known grade trends in the mineralization.

Figure 14-11: Directional semi-variograms for the short range high grade portion of the Main zone

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The modelled parameters defined from all semi-variograms are shown in Table 14-4.

Table 14-4: Semi-variogram model parameters

Hanging Wall Nugget 0.0694 Sill 0.175

Azimuth (°) Plunge (°) Range (°) Shape Direction1 228.00 8.00 50.05 Exponential Direction 2 326.52 46.50 15.70 Exponential Direction 3 130.65 42.39 10.22 Exponential Main Nugget 0.1029 Sill 0.1386

Azimuth (°) Plunge (°) Range (°) Shape Direction 1 228.00 8.00 35.06 Exponential Direction 2 326.52 46.50 22.76 Exponential Direction 3 130.65 42.39 4.88 Exponential Main Hi Nugget 0.025 Sill 0.073

Azimuth (°) Plunge (°) Range (°) Shape Direction 1 247.50 21.00 12.42 Exponential Direction 2 335.09 38.22 15.54 Exponential Direction 3 135.38 44.45 8.16 Exponential Footwall Nugget 0.645 Sill 0.186

Azimuth (°) Plunge (°) Range (°) Shape Direction 1 228.0 8.00 51.85 Exponential Direction 2 326.52 46.5 10.90 Exponential Direction 3 130.65 42.39 6.20 Exponential

All semi-variogram models were tested using MIK cross-validation routines to generate a file of original results, estimates and residual (Estimate – Actual Value) values.

MIK is a non-parametric Kriging method that does not rely on assumptions about the shape of the underlying data distribution and can therefore handle the modelling of spatially mixed data distributions. The probability that a data estimate exceeds a particular threshold (indicator) is estimated over a range of thresholds with the final estimate being an average of the estimated values. Threshold values were based on percentile values of the data distribution for each of the modelled zones (Table 14-5).

Table 14-5: Percentile values used for MIK thresholds

Percentile HW Main FW 1 0.03 0.14 0.01 5 0.06 0.27 0.05 10 0.09 0.32 0.07 20 0.12 0.45 0.10 25 0.14 0.51 0.12 30 0.15 0.59 0.14

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Percentile HW Main FW 40 0.19 0.74 0.15 50 0.22 0.90 0.18 60 0.27 1.13 0.20 70 0.34 1.44 0.26 75 0.38 1.66 0.30 80 0.45 1.97 0.35 85 0.55 2.40 0.45 90 0.73 3.22 0.60 95 1.21 4.89 0.96 97.5 2.15 6.69 1.51 99 4.19 10.40 2.40

The additional higher grade zone estimate, Main Hi, used the bins listed in Table 14-6.

Table 14-6: Bins and values used for MIK thresholds in the Main Hi zone

Bin Value 1 0.25 2 0.50 3 1.00 4 2.00 5 4.50 6 7.00 7 9.50 8 12.00

The output from the cross-validation routines for each of the semi-variogram models were used to construct QQ plots which compared the shape of the input distribution with the modelled data distribution. The results listed in Table 14-7 demonstrate the closeness of fit between the model estimates the original input values.

Table 14-7: Cross-validation results

Zone HW Main Main Hi FW Estimate Estimate Estimate Estimate Estimate Number of Points 4,782 6,247 691 1,588 Mean 0.42 1.54 4.7 0.32 Variance 0.22 0.86 4.41 0.13 Standard Deviation 0.47 0.93 2.10 0.36 Actual Composites Composites Composites Composites Number of Points 4,926 6,306 1,246 1665 Mean 0.44 1.54 4.44 0.32 Variance 1.58 4.83 13.13 0.37 Standard Deviation 1.26 2.2 3.62 0.61 CC 0.934 0.957 0.962 0.985 Rank CC 1.00 1.00 1.00 1.00 X Var / Y Var 7.28 5.63 2.98 2.94 Precision 189.94 86.92 38.34 83.33

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14.2.5 Block Model The block model was based on 25 m × 25 m × 10 m blocks rotated 45° around the Z axis to fit the blocks to the wireframe boundaries. The model extents and rotations are shown in Table 14-8 and Table 14-9.

Table 14-8: Model dimensions

Direction Model Minimum Block Size Number of blocks Model Maximum East 695800 25 60 697400 North 8453800 25 61 8455200 RL -25 10 50 465

Table 14-9: Model rotations

Rotation axis Rotation degrees Azimuth (Z) 45 Plunge (X) 0 Rotation (Y) 0

The blank block model was then coded to flag, blocks according to mineralised domain. When assigning the mineralization wireframes, blocks were sub-blocked to a minimum block size of 5 m × 5 m × 5 m. Backfilled and historical final mined surface and topographic surfaces together with oxidation surfaces were then assigned into the model. As with the ore zones, blocks were sub-blocked to ensure a good fit with the domains. Minimum block sizes in the model after all wireframes had been assigned are shown in Table 14-10.

Table 14-10: Model dimensions

Dimension Direction Minimum Maximum X NW–SE 5 25 Y SW–NE 5 25 Z Elevation 0.8333 10

There are 312 blocks, out of 310,000 blocks in the model, with a z dimension less than 1 m. These blocks occur adjacent to the subhorizontal backfill and historical final mined and topographic surfaces, base of soil and base of weathered wireframes.

For the current Mineral Resource estimate, the model used a copy of the model used and reported in 2017. The copy was then modified to include information on the blocks in the model between the backfilled surface and the historical final mined surface. This approach allowed check reports to be generated which confirmed the 2017 Mineral Resource estimate prior to reporting the current estimate.

For identification blocks in the model inside the various mineralised wireframes were coded HW, MAIN, FW and DYKE in a field called OWF_Code.

Blocks above the base of soil wireframe were coded SOIL in a field called SOIL_SAP; those above the base of weathering were coded SAP and the remaining blocks were assumed to be fresh rock and were coded FRESH.

Topography used in the 2017 model was coded BEL in a field called Topo_Code; blocks above that topography were discarded. Blocks below the backfill surface were coded BEL in a field called BF_Code, while those below the historical final mined surface were coded BEL in a field called Pit_Code.

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The SG values added to the 2017 model were retained in a field called SG_2017.

For this Mineral Resource estimate, new field called SG_2018 was added, this differed from the SG_2017 field as those blocks which were within the ore wireframes and were between the backfill and historical final mined surfaces were given a density of 2,000 kg/m3. The density values used are listed in Table 14-11.

Table 14-11: Assigned density values

Material 2017 2018 Soil 1,800 1,800 Saprolite 2,400 2,400 Fresh 2,730 2,730 Backfill 2,000 Note: Density values are in kg/m3, the modelling software used values in t/m3.

14.2.6 Interpolation The interpolation used a Multiple Indicator Kriging (MIK) algorithm, based on the median semi- variogram. Each of the three ore zones, HW, Main and FW, coded in the block model was interpolated separately, using the data points from inside that zone as the data source. The HW and FW zones were modelled using four interpolation runs. Blocks in the Main Zone were modelled using five interpolation runs, the first run in the Main Zone used the semi-variography developed using the high grade hits in the Main Zone (Main Hi). These were modelled using a small search ellipse. Modelling then proceeded as for the HW and FW zones. The search parameters for each interpolation run are listed in Table 14-12 and Table 14-13.

Table 14-12: Model parameters for interpolation runs other than Main Hi zone

Long axis Azimuth Plunge Intermediate axis Rotation Short axis Sectors 1 228 8 0.5 -46.5 0.1 8

Table 14-13: Model parameters for interpolation runs for Main Hi zone

Long axis Azimuth Plunge Intermediate axis Rotation Short axis Sectors 1 247 21 0.5 -46.5 0.1 8

The search ellipsoids for each interpolation run varied in size, the minimum number of data sources (drill holes) and points that could contribute to a block. The values used for each run are listed in Table 14-14.

Table 14-14: Search parameters for each interpolation run

Maximum Ellipsoid Minimum Minimum Maximum Minimum Zone Run points per length sources points points points sector FW1 1 35 2 1 10 10 4 FW2 2 70 2 1 10 10 2 FW3 3 140 2 1 10 10 2 FW4 4 280 1 1 10 10 2 Main 1 1 35 2 1 10 10 2 Main 2 2 35 2 1 10 10 4 Main 3 3 70 2 1 10 10 2 Main 4 4 140 2 1 10 10 2 Main 5 5 280 1 1 10 10 2

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Maximum Ellipsoid Minimum Minimum Maximum Minimum Zone Run points per length sources points points points sector HW1 1 35 2 1 10 10 4 HW2 2 70 2 1 10 10 2 HW3 3 140 2 1 10 10 2 HW4 4 280 1 1 10 10 2

In addition to a grade estimate for each block, information on the number of points and data sources, minimum and average distance to points contributing to the estimate, and kriging standard error was assigned to each block. Because an MIK algorithm was used, a grade estimate and the probability of the grade being in each of the grade bins (Table 14-5 and Table 14-6) was recorded. The grade for the highest grade bin for which the probability was 1 was then recorded as the grade estimate for that block.

In addition to the interpolation runs, a confidence indicator was given to each block depending on the interpolation run in which a block was estimated. Blocks estimated with the smallest radius and most data sources were given the highest ranking, while those interpolated using a subsequent interpolation run received a lower confidence ranking. Rankings ranged from 1 (most confident) to 5 (least confident). This information was subsequently used as a determinant when the model was classified for reporting.

Once constructed, the model was tested, using a QQ plot, against the source data to determine the relationship between the source and the modelled data, as shown in Figure 14-12.

Figure 14-12: QQ plot of the modelled data vs input data

As expected, the model shows that grades are smoothed; however, there is a high degree of correlation between the source data and modelled data up to a grade of 2.0 g/t Au, which suggests the model is a good fit to the bulk of the source data.

14.2.7 Model Classification and Reporting The classification of blocks into Measured, Indicated and Inferred Mineral Resource blocks used the following factors to determine the final block classification:

• Zoning of data based on data density (drill hole pierce points) as determined in an inclined long section

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• Block confidence as recorded from the interpolation run which estimated a block and average distance of points from the center of a block • Whether a block was located between the backfill and mined surfaces. Zoning The zoning of the data was determined from an inclined long section and the digitized boundaries over the section, based on the density of the drill holes. Holes in the Zone 1 were generally spaced 20 m apart; those in Zone 2 were generally 40 m apart and those in Zone 3 were 80 m apart. The zones are shown in Figure 14-13.

Figure 14-13: Ore classification boundaries used in the model

Block confidence and average distance Information on the confidence of the estimate, as determined by the interpolation run in which it was estimated, and the average distance to the data points used in estimating a block, was then combined and used to further limit the reporting classification given to a block. The interpolation runs used values of 1–4 to record the confidence in a block, with 4 being the highest confidence and 1 being the lowest. The average distance between the center of a block and the data points contributing to the block was recorded in the model. Blocks with a low confidence and which were estimated by points that on average were a long way from the block center were flagged and given a lower overall rating. This only affected blocks estimated in the last interpolation run for a mineralized domain. This information was then combined with the classification zone to produce the block classification used for reporting purposes in the 2016 and 2017 Mineral Resource estimates.

Blocks were classified for the 2016, 2017 and 2018 Mineral Resource estimates as shown in Table 14-15.

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Table 14-15: Block classification scheme

Classification Confidence Average distance Final Boundary colour Zone Indicator (m) classification Zone 3 Green =1 Not classified Zone 3 Green >1 Inferred Zone 2 Blue =1 Inferred Zone 2 Blue >1 Indicated Zone 1 Red =1 Indicated Zone 1 Red >1 >45 Indicated Zone 1 Red >1 <=45 Measured

This means that blocks interpolated in the last interpolation run and which fell in the Zone 3 of the long section were not given a classification. Other blocks in Zone 3 and which were interpolated in a run other than the last run were classified as Inferred Mineral Resource.

Blocks which were interpolated in the last interpolation run and which fell in Zone 2 of the long section were only given an Inferred classification. Other blocks in Zone 2 and which were interpolated in a run other than the last run were classified as Indicated Mineral Resource.

Blocks which fell in Zone 3 of the long section and which were interpolated in the last interpolation run or which had an average distance of more than 45 m were classified as Indicated Mineral Resource. Other blocks in Zone 1 which were interpolated in a run other than the last run or had an average distance of less than 45 m were classified as Measured Mineral Resource. Backfilled and Mined Surfaces The recognition that there were blocks in the model that were between the backfilled and historical final mined surfaces imposed a further constraint on the classification of blocks in the 2018 model. All blocks within the ore wireframes, and below the backfill surface and above the mined surface, were considered to be of a lower quality than other blocks. None of these blocks were therefore classified as Mineral Resource material.

14.2.8 Mineral Resource A Mineral Resource can only be declared for material which is considered to have potential for economic extraction at some point in the future. The cut-off at which a resource is reported should also meet this criterion, it should not include material which does not have reasonable potential to be mined and processed. The definition on a Mineral Reserve on the other hand applies a specific set of economic parameters to a Mineral Resource to determine which portions of the resource can be mined economically. In the case of the Posse deposit, economic modelling of the blocks in the model indicated that there are blocks with grades above 0.216 g/t Au that will be economic to mine. On this basis, the cut-off grade for the Mineral Resource has been set at 0.2 g/t Au. The Mineral Resource above a cut-off of 0.2 g/t Au declared for the Posse deposit is summarized in the Table 14-16 and Table 14-17, while a grade-tonnage curve for the deposit is shown in Figure 14-14.

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Table 14-16: Mineral Resource summary - 2018

Volume Tonnes Grade Metal % Category Above Domain Category (Mm3) (Mt) (g/t Au) (koz) Metal 0.20 1.6 4.4 0.69 97 17 Measured 0.20 2.4 6.4 0.80 170 23 HW Indicated 0.20 1.3 3.6 0.53 62 19 Inferred 0.20 2.4 6.6 2.1 440 79 Measured 0.20 3.6 9.8 1.6 500 71 MAIN Indicated 0.20 2.2 6.0 1.3 250 75 Inferred 0.20 0.50 1.3 0.40 17 3 Measured 0.20 0.88 2.4 0.52 40 6 FW Indicated 0.20 0.58 1.6 0.39 20 6 Inferred 0.20 4.6 12 1.4 560 100 Measured 0.20 6.8 19 1.2 710 100 ALL Indicated 0.20 4.1 11 0.92 330 100 Inferred Notes: HW = Hanging Wall MAIN = Main FW = Footwall. All figures have been rounded to two significant figures. A cut-off grade of 0.2 g/t Au has been used. Due to rounding, numbers may not sum correctly.

Table 14-17: Mineral Resource summary 2018 - Measured and Indicated only

Volume Tonnes Grade Metal % Category Above Domain Category (Mm3) (Mt) (g/t Au) (koz) Metal 0.20 1.6 4.4 0.69 97 17 Measured 0.20 2.4 6.4 0.80 170 23 HW Indicated 0.20 4.0 11 0.75 260 20 M&I 0.20 2.4 6.6 2.1 440 79 Measured 0.20 3.6 9.8 1.6 500 71 MAIN Indicated 0.20 6.0 16 1.8 940 72 M&I 0.20 0.50 1.3 0.40 17 3 Measured 0.20 0.88 2.4 0.52 40 6 FW Indicated 0.20 1.4 3.7 0.48 58 5 M&I 0.20 4.6 12 1.4 560 100 Measured 0.20 6.8 19 1.2 710 100 ALL Indicated 0.20 11 31 1.30 1,300 100 M&I Notes: HW = Hanging Wall MAIN = Main FW = Footwall. All figures have been rounded to two significant figures. A cut-off grade of 0.2 g/t Au has been used. Due to rounding, numbers may not sum correctly.

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Grade and Tonnage (Measured and Indicated) 2018 4 400

3.5 350

300 3

250 2.5 200 2 Tonnes (Mt)

Grade (g/tAu) 150

1.5 100

1 50

0.5 0 2.5 2.25 2.0 1.75 1.5 1.25 1.0 0.75 0.7 0.65 0.6 0.55 0.5 0.45 0.4 0.35 0.3 0.25 0.2 0.15 0.1 0 Cut-off (g/t) Grade Tonnes

Figure 14-14: Grade-tonnage curve Source: Posse, 2018.

The information in this Report that relates to the Mineral Resource estimates at Mara Rosa is based on information compiled by Gregory Keith Whitehouse, of the independent consulting firm Australian Exploration Field Services Pty Ltd (AEFS). Mr Whitehouse is a full-time employee of AEFS. Mineral Resource Modifying Factors The authors are not aware of any modifying factors such as mining methods, metallurgy, environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other factors which will have an impact on the Mineral Resource.

14.3 Interpretation and Conclusions The Posse deposit is hosted by a mylonitic shear hosted zone in a high greenschist to low amphibolite metamorphic terrain. The orebody strikes NE–SW and dips ~45° to the NW. On average, the orebody is ~30 m wide. Alteration is dominated by silicification, sericitisation, K-feldspar flooding and pyritisation. Gold is positively correlated with the intensity of silicification and total sulphide content and occurs as 10–100 µ sized particles along the margins of silicates and in association with pyrite

(FeS2) and frohbergite (FeTe2).

The new Mineral Resource estimated in this Report has established a Mineral Resource of ~31 Mt containing around 1.3 Moz Au at a grade of 1.30 g/t Au, above a cut-off grade of 0.20 g/t Au in the Measured Resource and Indicated Mineral Resource categories. A further 11 Mt containing 0.33 Moz Au at a grade of 0.92 g/t Au, above a cut-off of 0.20 g/t Au has been classified as Inferred Mineral Resource.

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The updated Mineral Resource includes Inferred, Indicated, and Measured Mineral Resource categories. Drilling completed in 2012 and reported as part of the 2016 Resource report significantly increased the confidence in the current Mineral Resource estimate. The differences between the current Mineral Resource and the 2011 Mineral Resource are shown in Table 14-18.

Table 14-18: Comparison between 2011 Mineral Resource and 2018 Mineral Resource

2011 percentage of 2018 percentage of Difference Category Mineral Resource Mineral Resource (koz) (%) (%) Measured 27 35 204 Indicated 61 44 -107 Inferred 11 20 174 Note: Due to rounding, numbers may not sum correctly.

Note that some of the increase in contained ounces is related to a change in cut-off grade used, from 0.5 g/t Au in 2011 to 0.2 g/t Au in this Report. The reduced cut-off grade is supported by the results of recent pit and mine optimization work.

The opinion of AEFS is that the character of the Mara Rosa Property, the Posse deposit and the Mineral Resource estimate reported herein is such that work to take the Posse Mineral Resource through an updated Pre-Feasibility and on to a Definitive Feasibility is recommended, together with ongoing exploration on the Property.

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15 Mineral Reserve Estimates This section presents the Mineral Reserve statement derived by SRK for the Project. This is constrained to a pit optimized by Whittle Consulting with the engineering design undertaken by SRK.

The Mineral Reserve is derived from the Measured and Indicated Mineral Resource presented in Section 14. The Mineral Reserve was optimized using a gold price of US$1,300/oz and the cost parameters outlined in Table 16-4.

The Mineral Reserve estimate is based on a diluted mining model with a variable cut-off grade. The variable cut-off grade is dependent on the grind size being used by the mill. Based on the result of the Enterprise Optimization undertaken by Whittle Consulting, the following two grind options were found to produce an optimal NPV:

• A P80 45 µm grind size which had a higher recovery of 92.0% Au, but a lower mill throughput rate and a higher milling energy consumption. The cut-off grade calculated for this grind size was 0.329 g/t Au

• A P80 75 µm grind size which had a lower recovery of 86.34% Au, but a higher mill throughput rate and a lower milling energy consumption. The cut-off grade calculated for this grind size was 0.216 g/t Au.

The mill is expected to undertake milling campaigns at each grind size to achieve the total production forecast.

The Whittle Consulting Enterprise Optimization has optimized the grind-throughput-recovery (GTR) of the mill over the LOM.

• The P80 45 µm grind size is used for the high grade component of the ore, with the aim of maximising the recovery and gold production

• The P80 75 µm grind size is used for the lower grade portion of the orebody, with the aim of maximising mill throughput while accepting a lower recovery from the lower grade ore.

The variable cut-off grade used over the LOM is shown in Figure 15-1.

AGL 1045A Cut-off Grade Assessment 0.450

0.400

0.350

0.300 off Grade (g/t) - 0.250

Au Cut 0.200

0.150 1 3 5 7 9 11 13 15 17 19 21 23 25 27 29 31 33 35 Period (qtrs)

Min Processed g/t Max Discarded g/t Implied cut-off grade g/t

Figure 15-1: LOM variable cut-off grade

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The estimation of Mineral Reserves used the following parameters:

• The Resource Model supplied by AEFS was diluted into a mining model using the following parameters: − Ore loss of 3% − Ore dilution of 3%, dilutive material grading 0.16 g/t Au − A variable cut-off grade based on the GTR of the mill • The Reserve Model assumes highly selective mining in the mineralized zones • An overall mining and haulage cost of US$2.04/t TMM • A processing cost of US$11.78/t for the 45 µm grind size, and a recovery of 92% and cut-off grade of 0.329 g/t Au • A processing cost of US$7.25/t for the 75 µm grind size, and a recovery of 86.34% and cut-off grade of 0.21 6g/t Au • An estimated selling cost of US$13.60/oz Au • Royalties of 6% of revenue. • A gold price of US$1,300/oz • A power cost of US$0.08/kWh • A diesel price of US$0.90 per litre.

Table 15-1: September 2018 Mineral Reserve estimate*

Diluted Diluted Contained Estimated Recoverable Mineral Reserve tonnes grade metal recovery metal (Mt dry) (g/t Au) (koz Au) (%Au) (koz Au) Proven 9.6 1.65 513 90.4% 464 Probable 14.2 1.26 574 90.8% 521 Total Mineral Reserve* 23.8 1.42 1,087* 90.6% 985 Note: *Open pit Mineral Reserves are reported at a variable cut-off grade dependent on GTR, assuming a metal price of US$1,300/oz, mining cost of US$2.04/t TMM, a variable processing cost dependent upon cut-off grade and GTR of Mill, an estimated selling cost of US$13.60/oz, a royalty of 6% of sales, and a variable processing recovery dependent on GTR. Mineral Resources reported are inclusive of all Mineral Reserves.

In SRK’s opinion, the technical parameters that form the basis of this Mineral Reserve statement are reasonable.

Other than discussed in this Report, SRK is not aware of any mining, metallurgical, infrastructure, permitting, environmental, legal, title, taxation, socio-economic, marketing or other relevant factors that could materially affect the Mineral Reserve estimate.

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16 Mining Methods 16.1 Geotechnical Design Parameters SRK has reviewed the ‘DFS Geotechnical Assessment 2013’ previously prepared by Coffey. In this PFS update, SRK has accepted the Coffey recommendations for the hanging wall gneiss, and the Northern and Southern schists. However, SRK has redesigned the Eastern Footwall in the amphibolite zone of the pit.

Table 16-1 summarizes the 2013 Coffey pit design recommendations. The Eastern pit sector has been divided into three design options, allowing for changes in design with variability of the foliation dip. The SRK-optimized design for the eastern wall has adopted the same approach, with footwall (FW) ‘blocks’ developed for the eastern wall, and aims to more closely follow the orebody for a steeper, more aggressive design.

Table 16-1: Coffey geotechnical recommendations 2013

Slope dip Design direction BFA BW BH IRSA Design IRSH OSH Domain Weathering Sector (°) (°) (m) (m) (°) Options (m) (m) From / To All All All Weathered 55 5 10 40 1 20 North 350/070 Fresh 65 6 20 52.5 1 Schist South 190/250 Fresh 65 6 20 52.5 1 65 14.5 20 40 1 Footwall 280 East 070/190 Fresh 65 10.5 20 45 2 260 Amphibolite 65 8.5 20 48 3 Hanging West 250/350 Fresh 65 6 20 52.5 1 wall Gneiss

The currently available geotechnical data is sparse, with ATV (acoustic televiewer) logging limited to six boreholes, of which only two are drilled perpendicular to the mineralized schist and continue into the FW. Figure 16-1 presents the geotechnical borehole database and the spatial layout within the pit. The structural data used in this assessment is the same dataset as used in the Coffey 2013 report.

Figure 16-1: Current geotechnical drilling with structural ATV disks shown

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16.1.1 Footwall Amphibolite The footwall model was developed through analyzing structural variability with depth. The structural database was filtered to include data from the boreholes that intercept the FW (MRP004, 010 and 013) allowing a strong foliation to be identified, dipping at 20°–40° to the northwest, which is coincident with the main orebody plunge. Figure 16-2 presents the stereographic projection of the discontinuities identified below the FW, and the spread in dip angle for the main FW foliation. A secondary foliation steeply dipping to the southeast was also identified; however, this is present in the base of MRP010 which is below the floor of the pit and is unlikely to be strongly present in the FW slopes. Planar sliding failure is expected to be the controlling instability mode.

Figure 16-2: Stereonet displaying structures below the footwall with strong NW dipping foliation, and dip variation for foliation presented in histogram

The Eastern pit sector has been further scrutinized by assessing orebody dip variability with depth, conducted by breaking the slope up into the following intervals, which are consistent with those used in previous PFS:

• Surface (approximately 440 m RL) to 320 m RL • 320 m RL to 260 m RL • 260 m RL to 200 m RL • 200 m RL to 100 m RL.

Vertical cross sections at 200 m spacing were used to identify zones of common dip for the orebody that were used to assign common slope properties, as shown in Figure 16-3 and Figure 16-4.

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Figure 16-3: East wall cross sections used for slope design

Figure 16-4: Example of orebody dip variability, addressed by developing different FW blocks for slope design

SRK adopted a variable inter-ramp angle (IRA) model to match the orebody plunge, represented by eight blocks. SRK performed kinematic stability analysis for the range of plunge within each domain to estimate the likely factor of safety (FoS) and potential bench failure volume. Empirical analyses have been conducted to confirm the spill berm widths required to adequately catch any failed material, with wider berms required with shallower bench face angles. The variable pit geometry elements are defined by a wireframe which can be used for design or optimization purposes. The pit design parameters based on the optimized geotechnical block model are presented in Table 16-2 and provide IRAs and Indicative Overall Slope Angles (IOSAs).

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Table 16-2: SRK-optimized Eastern wall pit design parameters

Geotechnical Blocks 1 2 3 4 5 6 7 8 Orebody dip (°) 46 48 53 40 51 46 46 47 BSA (°) 54.6 53.4 52.8 43.7 52.8 54.6 54.6 53.9 BSH (m) 100 100 100 100 100 100 100 100 BH (m) 20 20 20 20 20 20 20 20 BFA (°) 65 65 65 65 65 65 65 65 SBW (m) 6.1 6.9 7.3 14.5 7.3 6.1 6.1 6.6 GB (m) 15 15 15 15 15 15 15 15 IRA (°) 52.3 50.96 50.26 40 50.21 52.3 52.3 51.52 IOSA (°) 45.3 47.1 49.3 39.9 47.2 51.4 47.1 48.3

The base case bench geometry is:

• Bench height (BH) = 20 m • Bench face angle (BFA) = 65° • Spill berm width (SBW) - the previous Modified Ritchie approach required 8.5 m and in this case, SRK varied the dimension to match the orebody plunge between 6.1 m and 14.5 m. Wider berm widths are necessary where the orebody dip flattens • Bench stack height (BSH) i.e. the maximum unbroken inter-ramp height = 100 m separated by a geotechnical berm (GB) of 15 m.

Assumptions are:

• The block model intersections for >0.25 g/t Au show dips range from 40° to 51° and the amphibolite foliation dips between 30° and 40° in the same direction. For the PFS update, SRK has assumed that the footwall orebody plunge and foliation are coincident (the regional geological data suggests foliation dips in the 35° to 55° range, which would support this assumption); at the FS stage, it may be possible to refine this if more structural data is available. • Pit drainage (horizontal drains) are not needed and that the slopes are adequately depressurised and dewatered so that groundwater does not induce significant instability. • There is ~30 m of oxide on the FW side which will need to be checked in the FS design – it is expected that horizontal drains and berm drainage will be needed to prevent zones of erosion coalescing to create larger zones of failure and pit seepage which will increase water management costs. This design can be reassessed when accessing the old pits to identify suitable geometry. • Major structures need to be mapped in the old pit once access is re-established and used to develop a working structural geological model to assist pit design. • Standard ground control/ slope management procedures need to be adopted so that the design assumptions are validated during mining and the design is further optimized. Mapping of the footwall structures will be very important to maintain the optimal pit production, as well as checking for the potential for adverse footwall structures which could be unstable. • Good quality blasting of final walls and major intermediate cutbacks will be critical to good performance; pre-splitting (or similar blasting techniques) should be adopted.

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16.2 Mining

16.2.1 Introduction A PFS level open pit mining study has been completed, which consisted of adjustment of the Mineral Resource model, an Enterprise Optimization undertaken by Whittle Consulting, production scheduling, mining equipment and labour estimation, mining operating strategy, and mining cost adjustments from the previous PFS update. Capital and Operating costs have been estimated to within a precision range of ± 25%. No underground mining methods have been evaluated in this case.

16.2.2 Mining Model The Mineral Resource model supplied by AEFS was diluted into a Reserve Model using the following parameters:

• Ore loss of 3% • Ore dilution of 3%, dilutive material grading 0.16 g/t Au • Moisture content of 3% • A variable cut-off grade was used dependant on mill grind size:

− A P80 45 µm grind size which had a higher Au recovery of 92.0%. The break-even cut-off grade calculated for this grind size was 0.329 g/t Au

− A P80 75 µm grind size which had a lower Au recovery of 86.34%. The break-even cut-off grade calculated for this grind size was 0.216 g/t Au.

The Reserve Model assumes highly selective mining in the mineralized zones.

Table 16-3 shows the cut-off grade estimation for the 45 µm and 75 µm GTR cases.

Table 16-3: Estimation of cut-off grade

Realised GTR grind Mill GTR cost GTR Processing Cut-off Revenue gold price size throughput factor recovery cost grade (US$/g) (US$/oz) (µm) (tph) (%) (%) (US$/t) (g/t Au) 1,209 45 317.1 100.00 92.00 35.77 11.78 0.329 1,209 75 515.2 61.55 86.34 33.57 7.25 0.216

16.2.3 Pit Optimization Table 16-4 shows the parameters used for pit optimization, which was run at US$1,300/oz. The 0.85 Revenue pit shell (40) was chosen as the basis for Final Pit and Phase design.

Table 16-4: Pit optimization parameters

Parameters PFS Case Basis Included Mineral Resources Measured & Indicated only AEFS Total material movement 20 Mtpa Fleet Capacity Ore processed 2.5 Mtpa Plant Capacity Ore losses 3% SRK Ore dilution 3% at a grade of 0.16 g/t Au SRK Moisture content 3% SRK Cut-off grade Variable by GTR SRK Geotechnical By Domain Coffey & SRK

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Parameters PFS Case Basis Mining costs contractor Drilling (fresh material only) US$0.20/wet t SRK - PFS Mine Contractor Blasting (fresh material only) US$0.27/wet t SRK - PFS Mine Contractor Load/ Dump US$0.75/wet t SRK - PFS Mine Contractor Haulage time and distance Haul distance up to 500 m US$0.60/t SRK - PFS Mine Contractor Haul distance 501–1,000 m US$0.65/t SRK - PFS Mine Contractor Haul distance 1,001–1,500 m US$0.70/t SRK - PFS Mine Contractor Haul distance 1,501–2,000 m US$0.75/t SRK - PFS Mine Contractor Haul distance 2,001–3,000 m US$0.80/t SRK - PFS Mine Contractor Haul distance 3,001–4,000 m US$0.85/t SRK - PFS Mine Contractor Pit to crusher US$0.70/t SRK - PFS Mine Contractor Pit to North Dump US$0.75/t SRK - PFS Mine Contractor Pit to East Dump US$0.75/t SRK - PFS Mine Contractor Pit to North Dump - Soil US$0.80/t SRK - PFS Mine Contractor Pit to East Dump - Soil US$0.80/t SRK - PFS Mine Contractor Pit to stockpiles US$0.80/t SRK - PFS Mine Contractor Stockpiles to crusher US$0.75/t SRK - PFS Mine Contractor Stockpile rehandle load and dump US$0.75/t SRK - PFS Mine Contractor Processing cost (with 5% contingency) US$12.22/t Nominal 45 µm grind Processing Recovery Base case GTR – 45 µm 92.00% Amarillo testwork GTR – 75 µm 86.34% Amarillo testwork Revenue factors Gold price US$1,300/oz SRK PFS Update Refining, transportation, insurance, sales US$13.60/oz Amarillo Gold Price net of refining & transport US$1,286.40/oz Calculation Total royalties 6% Amarillo Net revenue US$1,209/oz Calculation Discount rate 5% SRK & Amarillo

16.2.4 Mine Layout The main infrastructure and facilities for the Project comprise the following:

• Internal roads and haul roads • Mine workshop and warehouses • Explosives and accessories magazines • Fuel facilities • Tire bay • Wash bay • Technical office • Core shed.

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The area selected for the mine facilities is close to the plant. The explosives and accessories magazines are to be located between the North and East Waste Dumps, at a minimum distance of 500 m. Figure 16-5 shows the proposed mine layout in 3D view.

Figure 16-5: Proposed mine layout

16.2.5 Life of Mine Plan The engineered ultimate pit design (including batter angle, berm widths and haulroads) and pit pushback phase designs have been designed in order to verify the technical feasibility of the optimal Whittle shells. The engineered pit designs are based on the selected US$1,300/oz Whittle pit shells.

Figure 16-6 shows the final pit design and its interaction with the orebody from several viewpoints.

Figure 16-7 shows the intermediate phase pushback pits and their relation to the final pit design.

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Figure 16-6: Final pit design and orebody (various views)

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Figure 16-7: Pit stages and final pit design

The engineered design includes two ramps in the hanging wall and two ramps in the footwall. Two-way ramps are to be used above 300 m RL, and one-way ramps below 300 m RL. Ramps in the pit are designed at a gradient of 10%.

Table 16-5 shows the ramp design criteria for the selected trucks.

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Table 16-5: Ramp design criteria

Parameter Value (m) Truck width 3.0 Ramp operation width (one-way) 6.0 Ramp operation width (two-way) 10.5 Truck tyre diameter 1.2 Safety berm height 0.6 Safety berm width 1.5 Ramp drainage width 1.5 Ramp total width (one way) 9.0 Ramp Total width (two way) 13.5

16.3 Production Scheduling The objective of the production scheduling was to meet the production needs of the mill and optimize NPV while maintaining good operational practices and a safely operated mine. Production scheduling involved definition of mining phases, the strategy for the utilisation of ore stockpiles and development of a mining schedule.

16.3.1 Schedule Assumptions The following assumptions were made in undertaking the production scheduling work:

• Only blocks classified as Measured Mineral Resource and Indicated Mineral Resource above the variable cut-off grade were scheduled and considered ore • Blocks classified as Inferred Mineral Resource and waste rock (below cut-off grade) were classified as waste • Pre-production in the year prior to the commencement of milling to ensure an initial supply of ore to mill. Pre-production TMM of 3.5 Mt, containing 109 kt of ore • Maximum TMM of 20 Mtpa • Operational benches: − Ore: 5 m − Waste: 10 m • Maximum annual vertical advance per phase: − Ore, one operational bench per month or 60 m per year − Waste, 7 operational benches per year or 70 m per year • Stockpiles: − High grade stockpile (>0.85 g/t Au) only used during pre-production, reclaimed during first year − Medium grade stockpile (0.5–0.7 g/t Au) − Low grade stockpile (0.216–0.5 g/t Au).

16.3.2 Schedule Results Table 16-6 and Table 16-7 shows the mining and processing schedules respectively.

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Table 16-6: Mining physicals

Mining Physicals LOM Totals Year 0 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Tonnes Moved Fresh Potential Ore 26,637,939 23,123 4,054,381 4,153,262 4,208,063 3,266,146 2,966,632 2,721,687 2,433,655 2,810,990 SAP Potential Ore 415,654 81,878 332,381 1,395 ------Soil Potential Ore 77,516 35,972 41,544 ------Fresh Waste 103,721,712 259,532 10,790,108 15,609,773 15,614,247 16,726,624 17,033,368 16,380,203 8,963,128 2,344,729 SAP Waste 5,006,243 1,663,172 2,943,989 214,207 177,645 7,230 - - - - Soil Waste 2,528,809 1,285,995 1,222,901 19,868 45 - - - - - Backfill Waste 773,132 156,942 614,696 1,495 ------Existing Stockpiles ------

Total Tonnes Mined 139,161,006 3,506,614 20,000,000 20,000,000 20,000,000 20,000,000 20,000,000 19,101,890 11,396,784 5,155,719 Tonnes Mined by Diluted Grade Band >= 0.85 g/t 15,368,607 45,590 2,458,268 2,041,793 2,018,389 1,957,405 1,455,399 1,645,551 1,575,976 2,170,236 >= 0.7g/t 1,613,568 6,901 153,935 228,186 358,726 193,129 221,637 164,061 66,433 220,561 >= 0.5 g/t 2,032,342 8,271 267,968 323,614 460,071 297,892 195,337 149,683 141,334 188,171 >= 0.38g/t 2,179,644 13,646 281,324 453,805 409,606 248,374 199,007 331,347 104,898 137,637 <0.38g/t 2,626,580 35,582 126,411 144,637 280,708 344,701 698,906 376,198 536,656 82,782

Total Tonnes Mined 23,820,741 109,990 3,287,906 3,192,035 3,527,499 3,041,500 2,770,286 2,666,840 2,425,297 2,799,387 Diluted Gold Grade (g/t) >= 0.85 g/t 1.93 1.45 2.13 2.40 1.70 2.02 1.54 1.65 1.70 2.00 >= 0.7g/t 0.78 0.80 0.77 0.77 0.77 0.78 0.78 0.79 0.78 0.78 >= 0.5 g/t 0.59 0.58 0.59 0.59 0.60 0.58 0.62 0.57 0.56 0.61 >= 0.38g/t 0.44 0.43 0.44 0.44 0.43 0.44 0.43 0.44 0.43 0.44 <0.38g/t 0.31 0.28 0.31 0.36 0.34 0.32 0.30 0.32 0.29 0.32

Average Diluted Gold Grade (g/t) 1.42 0.84 1.72 1.73 1.21 1.48 1.02 1.20 1.24 1.69 Strip Ratio (tonne Waste / tonne Ore)

Average 4.84 30.88 5.08 5.27 4.67 5.58 6.22 6.16 3.70 0.84

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Mining Physicals LOM Totals Year 0 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 ROM - Tonnes Plant - Direct Feed 23,270,898 - 3,012,977 3,106,708 3,447,902 3,041,500 2,770,286 2,666,840 2,425,297 2,799,387 Stockpiled 549,843 109,990 274,929 85,328 79,597 - - - - - Potential Ore Discarded to Topsoil Stockpile 12,373 6,471 5,902 ------Potential Ore Discarded to North Dump 3,065,445 - 1,133,816 962,036 680,467 197,943 87,957 3,225 - - Potential Ore Discarded to East Dump 232,551 24,512 683 585 97 26,703 108,389 51,621 8,358 11,603 Waste Discarded to Topsoil Stockpile 2,528,809 1,285,995 1,222,901 19,868 45 - - - - - Waste Discarded to North Dump 76,749,029 - 13,934,412 15,820,909 15,614,247 16,687,380 13,301,232 1,390,850 - - Waste Discarded to East Dump 32,752,059 2,079,646 414,381 4,566 177,645 46,474 3,732,137 14,989,353 8,963,128 2,344,729

Total ROM Movement 139,161,006 3,506,614 20,000,000 20,000,000 20,000,000 20,000,000 20,000,000 19,101,890 11,396,784 5,155,719 Stockpiles - Tonnes Plant - Direct from Existing Stockpiles ------Plant - From Prober Stockpiles 549,843 - 60,762 - - - 311,421 177,659 - -

Total Stockpile Movement 549,843 - 60,762 - - - 311,421 177,659 - -

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Total Tonnes Movement Ore Mined for Processing / Stkpl 25 4 2.0

20 3 1.5 15 2 1.0 10 Million Tonnes 5 Million Tonnes 1 0.5 Average Au Grade (g/t) Au Grade Average - - - PS Y02 Y04 Y06 Y08 Y10 PS Y02 Y04 Y06 Y08 Y10 Fresh Potential Ore SAP Potential Ore Soil Potential Ore >= 0.85 g/t >= 0.7g/t >= 0.5 g/t Fresh Waste SAP Waste Soil Waste >= 0.38g/t <0.38g/t Avg Grade Backfill Waste Mining Limit

Mining by Phase 10.0

5.0 Million BCM

0.0 PS Y02 Y04 Y06 Y08 Y10

Phase 1-0, Starter Phase 1-1, North Phase 1-2, South Phase 1-3, Wedge Phase 1-4, South 2 Phase 2-1, North Phase 2-2, South Phase 2-3, Wedge Phase 3 Phase 4-1, North Phase 4-2, South Phase 5-1, South Phase 6-1, South

Figure 16-8: Mining physicals

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Table 16-7: Processing physicals

Processing Physicals LOM Totals Year 0 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Tonnes ROM direct 23,270,898 - 3,012,977 3,106,708 3,447,902 3,041,500 2,770,286 2,666,840 2,425,297 2,799,387 Existing Stockpiles ------Prober Stockpiles 549,843 - 60,762 - - - 311,421 177,659 - - Total Tonnes 23,820,741 - 3,073,740 3,106,708 3,447,902 3,041,500 3,081,706 2,844,499 2,425,297 2,799,387 Average Diluted Grade (g/t Au) ROM direct 1.44 - 1.84 1.76 1.22 1.48 1.02 1.20 1.24 1.69 Existing Stockpiles ------Prober Stockpiles 0.54 - 1.26 - - - 0.53 0.31 - - Average Diluted Grade (g/t Au) 1.42 - 1.83 1.76 1.22 1.48 0.97 1.15 1.24 1.69 Tonnes Feed Base Case GTR – 45 µm 12,108,600 - 1,256,437 1,528,765 982,571 1,633,151 1,568,788 1,663,109 1,621,366 1,854,413 GTR4 – 75 µm 11,712,141 - 1,817,302 1,577,943 2,465,331 1,408,349 1,512,919 1,181,390 803,932 944,974 Total Tonnes Processed 23,820,741 - 3,073,740 3,106,708 3,447,902 3,041,500 3,081,706 2,844,499 2,425,297 2,799,387 Feed grade (g/t Au) Base Case GTR – 45 µm 2.10 - 2.74 2.81 2.33 2.22 1.48 1.64 1.67 2.16 GTR4 – 75 µm 0.72 - 1.20 0.75 0.78 0.62 0.44 0.45 0.37 0.76 Average Diluted Grade (g/t Au) 1.42 - 1.83 1.76 1.22 1.48 0.97 1.15 1.24 1.69 % Recovery by GTR used Base Case GTR – 45 µm 92.00% - 92.00% 92.00% 92.00% 92.00% 92.00% 92.00% 92.00% 92.00% GTR4 – 75 µm 86.34% - 86.34% 86.34% 86.34% 86.34% 86.34% 86.34% 86.34% 86.34% Overall Gold Ounces Fresh 973,325 - 152,843 159,858 121,363 131,622 87,099 93,879 88,429 138,231 SAP 9,844 - 8,280 24 - - 42 1,498 - - Soil 1,636 - 1,418 - - - 166 53 - - Total Ounces of Gold Produced 984,806 - 162,541 159,883 121,363 131,622 87,307 95,429 88,429 138,231

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Figure 16-9: Processing physicals

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Figure 16-10: Processing by grind size

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16.3.3 Mining Equipment and Operations The mine operation will use conventional open pit mining techniques and small size mining equipment to mine a total of 139 Mt of material (23.8 Mt of ore) over the LOM. Table 16-6 and Table 16-7 show the Project’s TMM.

The main mining equipment selected is of hydraulic excavators, 45 tonne on-highway trucks, and drills. Table 16-8 shows the equipment selection.

An overall availability of 80% has been assumed for the main production equipment and 75% for all other types. A maximum utilisation of 80% has been applied to the haulage trucks, 85% for loading equipment and 75% for drilling equipment. An efficiency factor of 83% (50 min/ 60 min) has been used to estimate productive hours for the main mining equipment.

Table 16-8: Mine equipment

Equipment Reference Model Excavator Liebherr R9100 Truck Scania G480 CB10x4 Drill Sandvik DP 1100i Bulldozer CAT D8T Grader CAT 140K Front end loader Liebherr L580 Crawler drill Atlas Copco ECM580 Rock breaker Atlas Copco H130S / CAT416E Crane Liebherr LTM 1030_2.1 Water truck Scania G440CB 6x4 NZ Blasting truck Scania G440CB 6x4 NZ Fuel & lube truck Scania G440CB 6x4 NZ Maintenance truck Scania P250DB 4x2 NZ Flatbed truck Scania G440CB 6x4 NZ Small backhoe CAT 416E Light vehicle Toyota Hilux 4 × 4 STR Lighting tower Terex RL4000 Dewatering system 200 HP

Supply storage and handling of explosive into the holes will be the responsibility of a specialised outsourced company, which will execute works under a service provision contract.

A blended explosive, 30% ANFO: 70% Emulsion, will be loaded into the holes. The blast plan design for both ore and waste is summarized in Table 16-9.

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Table 16-9: Blasting parameters

Blasting Parameter Unit Value Bench height m 10 Hole diameter mm 127 Hole diameter m 0.13 Burden m 3.0 Spacing m 4.5 Sub-drill m 1 Total hole depth m 11 Stemming m 1.95 Bottom charge m 0.9 Column charge m 8.15 Explosive density g/cm3 0.8 Specific charge kg/ml 9.8 Hole charge kg/hole 88.8 Volume m3 135 Explosive consumption kg/m3 0.66

16.3.4 Mining Labour Requirements Managers, administrative workers and some technicians will work day shift only. Equipment operators will either be permanent dayshift or on a rotating 3-shift basis depending on the type of equipment.

The workforce required varies between 146 and 342 workers throughout the operation of the mine. This includes allowances for holidays, sick leave and training.

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17 Recovery Methods The authors are not aware of any metallurgical study work which has been undertaken since the effective date of the 2016 Report (21 July 2016). Work prior to that date was covered in the 2016 Report.

However, in the Enterprise Optimization undertaken by Whittle Consulting a strategy of grind size variance was implemented.

17.1 Variance of Grind Size The ultimate grind size selected for the mill effects both mill throughput, power consumption and gold recovery. Several Grind-Throughput-Recovery (GTR) scenarios have been evaluated, two were chosen and modelled for the Project.

• A P80 45 µm grind size which had a higher recovery of 92.0% Au, but a lower mill throughput rate and a higher milling energy consumption

• A P80 75 µm grind size which had a lower recovery of 86.3% Au, but a higher mill throughput rate and a lower milling energy consumption.

The Enterprise Optimization has optimized the GTR of the mill over the LOM. Table 16-7 shows the GTR strategy used.

• The P80 45 µm grind size is used for the high-grade component of the ore, with the aim of maximising the recovery and gold production

• The P80 75 µm grind size is used for the lower grade portion of the orebody, with the aim of maximising mill throughput while accepting a lower recovery from the lower grade ore.

SRK understand that Amarillo will undertake milling campaigns of each grind size throughout the LOM.

Costs and throughputs were proportioned based on the power differential which is based on the efficiency factors and calculations from Amarillo.

Amarillo is conducting ongoing Bond Work Index (BWI) testwork that will give guidance on the grind power demand.

The power demand for the nominal grind is a key element in the GTR model. Ongoing testwork is recommended.

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18 Project Infrastructure The authors are not aware of any additional study work pertaining to the infrastructure which has been undertaken since the effective date of the 2011 Report. Work prior to that date was covered in the 2011 Report.

Power will be supplied by a 64 km 138 kV power supply line and substation in the nearby city of Porangatu.

Water requirements will be largely supplied by return water from the tailings dam and from dewatering the pit with makeup quantities coming from wells (for good quality water used to supply the potable water plant and for the makeup of reagents and pump seal water) as well as water from the local river during the wet season.

Lime will be sourced locally, cyanide in solid form from the Candeias plant in the state of Bahia, and mill balls will be imported from Chile.

Other infrastructure including an operations camp, surface workshops and warehouse, canteen and administration buildings have been developed to support cost estimation and development of general arrangements for the Project.

The tailings storage facility (TSF) was updated in the 2017 PFS update and no further updates have been undertaken. The system of tailings disposal will consist of a tailings dam situated southwest of the industrial plant occupying 140 ha. Figure 18-1 shows the general layout of the TSF.

The design for the TSF aims to optimize tailings storage capacity by maximising tailings density and reducing environmental and societal impact.

The construction of the TSF will be undertaken in four stages, including an initial phase followed by three successive raising of dam walls during the anticipated 8-year LOM. The timing of the lifts was defined on the anticipated risings of reject from the process plant during the LOM.

The TSF design is based on the following parameters:

• Total production of plant tailings (reject) of 19 Mt • Estimated ROM production of 2.5 Mtpa • Tailings slurry density nominally 59% solids (by weight) • Deposited tailings dry density has been assumed to be 1.40 t/m³ • Estimated tailings beach slope of 1%.

Table 18-1 presents the anticipated construction phases showing the evolution of TSF capacity with time and increasing dam crest height.

Table 18-1: Staged storage capacity, quantities and implementation

Dry mass Crest elevation Implementation Phase stored by phase (m RL) (years after start) (tonnes)

1 3,876,875 468 Start up 2 3,344,750 473 2 3 4,885,232 478 4 4 4,907,956 482 7

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SRK notes that there has been an increase of 4.8 Mt of ore being processed during the life of the plant, from 19 Mt to 23.8 Mt. A final tailings dam lift may be required in Year 7 of operations at an estimated cost of US$4.61M. At this stage, the additional sustaining capital has not been included in the capital costs and financial modelling undertaken in Section 21 and 22. SRK does not consider the economic impact of this potential additional sustaining capital would be material to the estimation of the Mineral Reserve.

Figure 18-1: General layout of the TSF

The dam wall will be constructed using a combination of soil and rock. The internal dam wall slopes will have an inclination with the horizontal component twice the vertical component or 2(h):1(v) and covered with an outer layer of compacted clay derived during pre-stripping or borrow areas. The walls and internal base of the TSF will be rendered impermeable with a 1 mm thick geomembrane (HDPE or equivalent material). An internal dam wall drainage system composed of a filter using sand and

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crushed stone is included adjacent to the clay liner. The drainage system will be capable of capturing any percolating water and will act as transition material to the principal and larger volume of rock and large waste blocks from the mine pit. The downstream slope will also have a horizontal component twice the vertical component or 2(h):1(v).

A thickened slurry of tailings with 59% solids will be discharged subaerially onto a beach from a number of spigots located along the main embankment crest.

The following additional investigations are recommended:

• Geotechnical investigation for the TSF • Drilling and equipping of the monitoring bores • Drilling and permeability testing of investigation bores • Laboratory testing of tailings and waste rock geochemistry and geotechnical properties • Confirmation of geotechnical parameters for the embankment construction materials.

The next stage of design work is a comprehensive feasibility design report for the TSF.

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19 Market Studies and Contracts The Project will produce gold bars containing about 95% gold. These bars will be refined to produce 99.9% purity gold and the refined gold will be sold to banks or other financial institutions either in Brazil or offshore on a spot price basis to capture the highest price.

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20 Environmental Studies, Permitting, and Social or Community Impact The authors are not aware of any further environmental, permitting, social or community study work that has been undertaken since the effective date of the 2011 Report. Work prior to that date was covered in the 2011 Report and is summarized below.

20.1 Physical and Natural Environment The climate of the Mara Rosa region is characterised by a hot, wet summer and a cool, very dry winter. Average temperatures in May/June are a pleasant 24°C, whereas average temperatures in August/ September reach 28°C. Average rainfall as measured at Estrela de Norte (30 km to the north) is 1,679 mm (1971–2010). Maximum humidity over 80% is reached in April and December, while minimum humidity of ~40% is experienced in July and August.

Studies of the natural environment were undertaken to establish baseline conditions for flora, fauna and the aquatic environment at the Mara Rosa site for use in identifying possible impacts from planned gold mining operations.

The Project is situated within the Cerrado Biome, an immense tropical savannah that constitutes Brazil’s second largest plant formation found in the Central Plateau regions covering some 23% of the country. This biome is one of a select few on Earth displaying both high biodiversity and threatened ecological status. The high human population of the area has resulted from socioeconomic intervention intended to transform the region into the country’s breadbasket.

Satellite imagery has indicated four distinct vegetative communities within the Project area and biological studies have been undertaken for each community.

The vegetation survey has evaluated any plant formations of special interest, and their preservation. The ecological functions and the environmental services of the vegetation were studied to ascertain any species protected by law. The aim of the study has been to develop a baseline database of information for the mining company to plan a sustainable method of mining and help develop mitigation measures and/ or create conservation areas where preservation or recovery will be implemented. This process must be undertaken in accordance with Ordinances 14/2001 and 15/2001.

The studied fauna of the area comprises some 16 mammal species, 22 reptile and amphibian species, and 79 bird species. Of the mammal species, two are considered vulnerable to extinction and will require further study. Although none of the reptile or amphibian species are vulnerable, the inventory data may be too few and inadequate for population estimation and conservation status. The bird species are highly dependent on the forest environments and will require careful monitoring of population levels so that rapid environmental changes caused by human activity do not adversely affect avifauna communities.

The aquatic environment includes three sub-basins, namely the Upstream Basin for the Ouro River; the sub-basin for Lambari Creek and the sub-basin for the Antas River.

Surface water quality in the old mine pits is satisfactory and the pit water can be safely discharged. Study of the phytoplankton and zooplankton communities showed a relatively high taxonomic complexity and a typical tropical freshwater state for retained water.

The benthic macro-invertebrate community was found to be extremely poor and draining the pits will therefore not cause problems of general diversity.

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Heavy metal concentrations were identified in sediments from the pits and it is recommended that sediment at the bottom of these pits should be transferred and impounded in the rejects dam, thus reducing the possibility of contaminating local surface water courses.

20.2 Hydrology and Hydrogeology Mining was last carried out at Mara Rosa in 1995. Significant environmental changes from the natural state include partially re-vegetated waste and tailings heaps and deforestation, but natural regrowth has hidden the worst of the degradation. The two old mining pits are now water-filled.

Using the updated Master Plan (site layout), a plan for the monitoring and sampling of surface and groundwater has been developed that will allow interpretation of water flows and quality for compliance with legal requirements and permitting.

Mara Rosa town, 5 km from the mine site, is supplied with water from a strong aquifer, but the bedrock geology in the planned mine site is not favourable for adequate groundwater supplies.

The paucity of seasonal hydrological data for the study area has precluded detailed estimates of available surface water in the Ouro River basin. However, preliminary estimates suggest that even after required minimum flow and allowances for agricultural purposes, some 626 m3/h may be available; this is far in excess of the projected needs of the Project (114 m3/h).

Previous surface and groundwater chemical analyses have not detected any contaminant values that exceed CONAMA Class II purity limits.

Recommendations are made to protect existing monitoring wells from airborne contamination. The old tailings from a different orebody (Zacarias) in the Baribras area may present an acid rock drainage risk which should be investigated and a mitigation plan implemented.

Continuing monitoring will be required within the mine area for diagnosis, assessment and mitigation plans for environmental damage especially adjacent to the Amarillo Posse project areas bordering streams and rivers.

20.3 Social Environment A preliminary socio-economic study of the demographics, land use, production and economics and quality of life has been undertaken using standard indicators and both direct and indirect survey methods. The economic base of Mara Rosa is small scale agriculture and livestock. The area is well serviced for a broad range of normal social and community facilities and is considered to have a high human development index in comparison with other areas of Brazil.

Future mining operations are viewed favourably due to the generation of employment and income. However, these operations will require necessary improvements to the physical and social infrastructure, and licensing will be required. The environmental licensing will require a more detailed archaeological survey to be completed.

20.4 Waste and Tailings Disposal, Monitoring and Water Management The tailings basins will be constructed with impermeable floors (clay followed by HDPE) as per mandates from State Environmental Agency, SECIMA.

Daily water consumption of 5,000 m3 will be largely supplied from pit dewatering and tailings dam return water. Supplementary water is expected to be largely surface water from streams for which a more detailed hydrological study will be required.

Air and water quality will be indicators to monitor during the LOM. The objective to restrict human intervention outside the service area must also be monitored through indicative studies.

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20.5 Permitting There are numerous laws and regulations governing mining and environmental issues at the federal, state and municipal levels of authority in Brazil. Amarillo has engaged a local Brazilian Environmental company to undertake all environmental permitting requirements for the Project.

An EIA-RIMA has been submitted to the relevant environmental authority (after 2011) and following a public audience in February 2016, the LP (provisional environmental licence) was obtained in May 2016.

SRK understands a further licence called the LI is required; this can only be applied for after more comprehensive study information has been obtained, including basic engineering to specify processing plant parameters more accurately.

20.6 Mine Closure A fundamental part of the permitting process is the Mine Closure Plan. A general view of requirements in this regard is in the process of being formulated for permitting purposes, and includes:

• Dismantling and selling or removing equipment for scrap, burying foundations, covering with soil and revegetating • TSF closure through capping, soil cover and revegetating • Contouring, covering and revegetating waste rock dumps • Passive flooding of the mine pits with perimeter fencing installation and water quality monitoring.

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21 Capital and Operating Costs 21.1 Approach A financial model has been developed to assess the operating costs and capital expenditure expected for the Project. The Project is to be based on a contract mining scenario as determined by Amarillo.

SRK used the project costs outlined in the previous 2017 PFS Update undertaken by SRKBR. Costs in Brazil have been affected by the recent devaluation of the Brazilian real, from R$3.2 to R$3.6 to the US$. As a result of this exchange rate shift, the estimated Brazilian component of costs was re- estimated in US$ terms by the project team (SRK-Whittle Consulting-Amarillo).

During any future DFS, a complete review of all operating and capital costs assumptions will be necessary to ensure a DFS level of estimation accuracy. All major equipment should be re-quoted and new contractor quotes obtained for any future DFS process.

21.2 Operating Costs The variable mining costs unit rates used in the cost model are shown in Table 21-1.

Table 21-1: Mining variable operating costs

Parameters PFS Case Basis Mining Costs Contractor Drilling (Fresh material type only) US$0.19/t wet t SRK PFS adjusted by exchange rate Blasting (Fresh material type only) US$0.27/wet t SRK PFS adjusted by exchange rate Load/ Dump US$0.73/wet t SRK PFS adjusted by exchange rate Haulage time and distance Haul distance up to 500 m US$0.58/t SRK PFS adjusted by exchange rate Haul distance 501–1,000 m US$0.63/t SRK PFS adjusted by exchange rate Haul distance 1,001–1,500 m US$0.68/t SRK PFS adjusted by exchange rate Haul distance 1,501–2,000 m US$0.73/t SRK PFS adjusted by exchange rate Haul distance 2,001–3,000 m US$0.78/t SRK PFS adjusted by exchange rate Haul distance 3,001–4,000 m US$0.83/t SRK PFS adjusted by exchange rate Pit to crusher US$0.68/t SRK PFS adjusted by exchange rate Pit to North Dump US$0.73/t SRK PFS adjusted by exchange rate Pit to East Dump US$0.73/t SRK PFS adjusted by exchange rate Pit to North Dump - Soil US$0.78/t SRK PFS adjusted by exchange rate Pit to East Dump - Soil US$0.78/t SRK PFS adjusted by exchange rate Pit to Stockpiles US$0.78/t SRK PFS adjusted by exchange rate Stockpiles to Crusher US$0.73/t SRK PFS adjusted by exchange rate Stockpile rehandle load and dump US$0.73/t SRK PFS adjusted by exchange rate

The fixed period costs used in the cost model are shown in Table 21-2.

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Table 21-2: Estimate of fixed costs

Parameters PFS Case (US$M/yr) Basis Mining - Fixed Cost $1.65 SRK-Amarillo-Whittle Processing - Fixed Cost $7.35 SRK-Amarillo-Whittle G&A $5.55 SRK-Amarillo-Whittle Total $14.55

Table 21-3 presents a summary of the processing costs used in the cost model.

Table 21-3: Processing costs summary

Parameters PFS Case Basis Processing Cost Variable GTR 45 µm US$6.69/t SRK adjusted by exchange rate Processing Cost Fixed GTR 45 µm US$4.53/t SRK adjusted by exchange rate Processing Cost (5% Contingency) US$0.56/t SRK adjusted by exchange rate Total Processing Cost GTR 45 µm US$11.78/t SRK adjusted by exchange rate Processing Recovery GTR – 45 µm 92% Au Amarillo testwork Processing Cost Variable GTR 75 µm US$4.12/t SRK adjusted by exchange rate Processing Cost Fixed GTR 75 µm US$2.79/t SRK adjusted by exchange rate Processing Cost (5% Contingency) US$0.34/t SRK adjusted by exchange rate Total Processing Cost GTR 75 µm US$7.25/t SRK adjusted by exchange rate Processing Recovery GTR – 75 µm 86.3% Au Amarillo testwork

Table 21-4 shows the estimated cash cost over the LOM for a total production of 984.8 koz.

Table 21-4: LOM cash cost estimate

LOM Cash Cost Estimate Total Cost (US$M) Unit Cost (US$/oz) Operating Cost Estimate Mining & Processing Variable $417.8 $424.3 Mining & Processing Fixed $73.7 $74.8 Site G&A $45.4 $46.1 Operating Cost $536.9 $545.2 Adjusted Operating Costs Estimate Refining, Transportation, Insurance $13.4 $13.6 Royalties $73.5 $74.7 Adjusted Operating Costs $623.8 $633.5 All-in Sustaining Cost (AISC) Estimate Sustaining Capital $17.4 $17.7 Reclamation Costs $4.0 $4.1 AISC $645.2 $655.3

21.3 Capital Costs The total forecast capital expenditure over the LOM is US$140.3M. This consists of US$122.9M of pre-production capital to bring the mine into operation, and US$17.4M of sustaining capital over the LOM. The summarized capital expenditure is shown in Table 21-5.

A summary of the LOM cost model is shown in Table 21-6.

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Table 21-5: Summary of capital expenditure

Parameters PFS Case (US$M) Basis Mine Contractor $5.3 SRK - PFS Processing Capital $104.3 Amarillo-adjusted by exchange rate Tailings Dam $23.0 SRK - PFS Working Capital $7.7 SRK - PFS Total Capital $140.3

Salvage - contractor -$12.1 SRK - PFS Reclamation $4.0 SRK - PFS Total Capital - Salvage + Reclamation $132.2

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Table 21-6: LOM operating and capital cost summary

LOM Year 0 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Operating and Capital Cost Summary (US$M) (US$M) (US$M) (US$M) (US$M) (US$M) (US$M) (US$M) (US$M) (US$M) Variable Costs Mining Costs - Drilling Costs - Ore 4.78 0.02 0.66 0.64 0.71 0.61 0.56 0.54 0.49 0.56 Mining Costs - Drilling Costs - Waste 21.03 0.05 2.25 3.17 3.18 3.38 3.42 3.30 1.80 0.47 Mining Costs - Blasting Costs - Ore 6.53 0.03 0.90 0.87 0.97 0.83 0.76 0.73 0.66 0.77 Mining Costs - Blasting Costs - Waste 28.71 0.07 3.07 4.33 4.34 4.61 4.67 4.50 2.46 0.65 Mining Costs - Load and Dump - Ore 17.93 0.08 2.48 2.40 2.66 2.29 2.09 2.01 1.83 2.11 Mining Costs - Load and Dump - Waste 86.83 2.56 12.58 12.65 12.40 12.77 12.97 12.37 6.75 1.77 Mining Costs - Haul to Plant 18.06 0.00 2.18 2.35 2.61 2.33 2.21 2.12 1.95 2.31 Mining Costs - Haul to Stockpiles 0.44 0.08 0.22 0.07 0.07 0.00 0.00 0.00 0.00 0.00 Mining Costs - Haul to North Dump 57.83 0.84 10.98 11.65 11.71 12.29 9.38 0.98 0.00 0.00 Mining Costs - Haul to East Dump 26.13 1.57 0.30 0.00 0.13 0.06 3.05 11.75 7.26 2.01 Rehandling Cost for Stockpiles to Plant Load and Dump 0.41 0.00 0.05 0.00 0.00 0.00 0.23 0.13 0.00 0.00 Rehandling Cost for Stockpiles to Plant Haulage 0.41 0.00 0.05 0.00 0.00 0.00 0.23 0.13 0.00 0.00 Processing - Crushing cost 6.18 0.00 0.80 0.81 0.89 0.79 0.80 0.74 0.63 0.73 Processing - Mill power cost 42.19 0.00 5.19 5.46 5.46 5.46 5.46 5.22 4.62 5.32 Processing - Mill non-power cost 38.00 0.00 4.67 4.92 4.92 4.92 4.92 4.70 4.16 4.79 Processing - CIL cost 62.37 0.00 8.05 8.13 9.03 7.96 8.07 7.45 6.35 7.33 Total Variable Costs 417.84 5.32 54.41 57.46 59.08 58.31 58.83 56.67 38.96 28.82 Fixed Costs and Capex Total Mining Period Costs 14.86 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 Total Processing Period Cost w/ 5% contingency 58.80 0.00 7.35 7.35 7.35 7.35 7.35 7.35 7.35 7.35 G&A w/ 5% contingency 45.38 1.00 5.55 5.55 5.55 5.55 5.55 5.55 5.55 5.55 Pre-production Capex 122.89 122.89 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 Sustaining Capex 17.40 0.00 0.93 4.44 4.52 4.61 0.00 0.00 0.00 2.90 Salvage -12.08 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 -12.08 Reclamation 4.00 0.00 0.00 0.00 0.35 0.35 0.35 0.35 0.35 2.25 Total 251.25 125.54 15.47 18.99 19.42 19.51 14.90 14.90 14.90 7.62

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22 Economic Analysis 22.1 Introduction The overall economics of the Project have been evaluated using conventional Discounted Cash flow (DCF) methods. The production schedules, capital expenditures and operating costs have previously been discussed in this Report. The following key parameters were integral to the construction of the cashflow model and the economic results:

• A Base Case gold price of US$1,300/oz was used • The analysis was based on 100% equity financing with no debt component • All revenues and costs are reported in ‘Real’ constant US$ terms without escalation.

The economic analysis presented in this section contains forward-looking information with regards to the Mineral Reserve estimates, commodity prices, exchange rates, proposed mine production plan, projected recovery rates and processing costs, infrastructure construction costs and schedule. The results of the economic analysis and are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.

SRK’s economic analysis is for the purpose for Mineral Reserve estimates only.

22.2 Cashflow Assumptions

22.2.1 Mine Production Sequence Table 16-6 and Table 16-7 shown previously in this Report outline the proposed Mining and Processing Schedules.

22.2.2 Operating and Capital Costs The operating and capital costs used in the financial model are shown in Table 21-6.

22.2.3 Metallurgical Recovery The two metallurgical recoveries were used in the modelling and they are as follows:

• P80 45 µm grind size: 92% Au recovery

• P80 75 µm grind size: 86.3% Au recovery • Average LOM Recovery: 90.6% Au recovery.

22.2.4 Metal Prices and Net Revenue The long-term gold price incorporated into the cashflow model was US$1,300/oz. Table 22-1 shows the calculation of net revenue in US$/oz, after adjustment for refining charges and royalties.

Table 22-1: Net revenue calculation

Revenue Factors Units PFS Case Basis Gold price US$/oz $1,300 SRK PFS Update Refining, transportation, insurance, sales US$/oz $13.60 Amarillo Gold price net of refining & transport costs US$/oz $1,286.40 Calculation Total royalties % 6% Amarillo Net revenue per oz US$/oz $1,209 Calculation

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22.2.5 Salvage Value The salvage value estimate of US$12.08M from the previous SRKBR 2017 PFS update was used in the financial and cost modelling.

22.2.6 Taxes A Brazilian income tax rate of 25% was used. Provision for the 9% Brazilian Social Contribution Tax was also made, bringing the effective tax rate to 34%.

22.2.7 Working Capital A Working Capital amount of US$7.73M has been included into Year 0 of the Financial Model. This provides approximately two months of operating cost requirements.

22.2.8 Reclamation The mine reclamation cost US$4.0M from the previous SRKBR 2017 PFS update was used in the financial and cost modelling.

22.3 Discounted Cashflow Result Table 22-2 is a summary of the LOM financial model outputs.

Table 22-2: LOM physicals

LOM physicals Units Value Material movement LOM Total material movement (TMM) Mt 139.16 Total waste movement Mt 115.34 Total ore mined Mt 23.82 Average ore grade g/t 1.42 Processing LOM Ore processed Mt 23.82 Ore grade processed g/t 1.42 Average recovery % 90.59% Gold produced koz 985

The LOM financials from the financial model are shown in Table 22-3. The resultant internal rate of return (IRR) is 50.8%

Table 22-3: LOM financials

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The results of the DCF analysis is shown in Table 22-4. The IRR produced is 50.8%.

Table 22-4: DCF result

Total Result (US$M) NPV (post-tax @ 5% discounted rate) 244.3 NPV (post-tax @ 7.5% discounted rate) 212.7 NPV (post-tax @ 10% discounted rate) 185.4

LOM Costs and Revenues are shown below in Figure 22-1.

Figure 22-1: LOM cashflow and margin Note: Run 045A – reference to optimization scenario number.

22.4 Sensitivity Analysis SRK undertook a high-level sensitivity analysis on the DCF Model, the results of which are shown in Table 22-5 and Figure 22-2.

The Project is most sensitive revenue, and least sensitive to capital expenditure.

Table 22-5: Sensitivity analysis result

Variance OPEX (US$) CAPEX (US$) Revenue (US$) 25% 170.3 209.2 406.1 20% 185.6 216.2 373.7 15% 200.3 223.3 341.4 10% 215.0 230.3 309.0 5% 229.6 237.3 276.7 0% 244.3 244.3 244.3 -5% 259.0 251.3 212.0 -10% 273.7 258.4 179.6 -15% 288.3 265.4 147.3 -20% 303.0 272.4 113.5 -25% 317.7 279.4 79.5

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Sensitivity $450.0 $400.0 $350.0 $300.0 $250.0 $200.0 $150.0 $100.0 $50.0 $0.0 -30% -20% -10% 0% 10% 20% 30%

OPEX NPV US$M @ 5% CAPEX NPV US$M @ 5% Revenue NPV US$M @ 5%

Figure 22-2: Sensitivity spider chart Table 22-6 shows the sensitivity of various financial parameters to gold price.

Table 22-6: Sensitivity to Revenue

Au price US$/oz $1,000 $1,100 $1,200 $1,300 $1,400 $1,500 IRR % 23% 33% 42% 51% 59% 68% Payback Period Years 4.2 3.3 2.8 2.6 2.4 2.3

NPV0% (After Tax) US$M 138 202 263 324 386 447

NPV5% (After Tax) US$M 91 144 194 244 295 345

NPV7.5% (After Tax) US$M 73 121 167 213 259 305

NPV10% (After Tax) US$M 57 101 143 185 228 270

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22.5 Comparison to Previous Studies Table 22-7 compares the 2018 PFS Update by SRK with the two previous PFS reports – Coffey (2011) and SRKBR (2017).

Table 22-7: Comparison to previous studies

2011 Coffey PFS 2017 SRKBR PFS 2018 SRK PFS Category Units (Owner Operator) (Contractor Mine) (Contractor Mine) Exchange rate US$/BRL 1.9 3.2 3.6 Initial capital US$M 183.6 132.3 122.9 Sustaining capital US$M 11.4 16.5 17.4 Total LOM capital US$M 195 148.8 140.3 Post-tax NPV @ 5% US$M 178.5 178.3 244.3 Post-tax IRR % 26.6 35.2 50.8 Cash operating cost US$/oz 464 545 545 (excluding royalty & refining) Cash operating cost US$/oz 524 603 633 (including royalty & refining) AISC US$/oz 546 627 655 (including sustaining capital & closure) Tonnes of ore processed Mt 17.1 19 23.8 Grade of ore processed g/t 1.72 1.63 1.42 LOM strip ratio (waste: ore) t:t 8.2: 1 4.5: 1 4.84: 1 Resources Measured & Indicated koz 1,174 1,260 1,300 Resources cut-off grade g/t 0.50 0.35 0.20 Reserves Proven & Probable koz 945 998 1,087 Reserves cut-off grade g/t 0.5 0.38 Variable US$1,300/oz 0.85 US$1,100/oz Pit US$1,200/oz Pit Revenue Factor Shell Shell Gold price US$/oz Pit Shell US$1,200/oz US$1,200/oz US$1,300/oz Financials Financials Financials 3% loss & 3% 3% loss & 3% 3% loss & 3% Mining loss & dilution % dilution dilution dilution Metallurgical recovery % 92 92 Variable

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23 Adjacent Properties Several other significant mineral deposits occur in the Mara Rosa region including the Zacarias gold- silver-barite deposit and the Chapada copper-gold deposit, in addition to numerous historical prospects and garimpos.

A detailed discussion of these various properties was provided in Hoogvliet Contract Services & Australian Exploration Field Services Pty Ltd (2011) that is filed at sedar.com and to which the reader is referred.

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24 Other Relevant Data and Information The authors are not aware of any additional information relating to the mineralization at the Project which is of a material nature.

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25 Interpretation and Conclusions The economic analysis presented undertaken contains forward-looking information with regards to the Mineral Reserve estimates, commodity prices, exchange rates, proposed mine production plan, projected recovery rates and processing costs, infrastructure construction costs and schedule. The results of the economic analysis are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.

SRK’s economic analysis is for the purpose for Mineral Reserve estimates only.

The economic analysis performed from the results of the PFS demonstrates an economic viability of the Project using the assumptions considered. In SRK’s opinion, the Mineral Reserves declared are valid.

On the basis of SRK’s analysis, the Project generates a post-tax IRR of 50.8%, a post-tax net present value (NPV5%) of US$244.3M, and post-tax project payback of 2.5 years based on a gold price of US$1,300/oz and a US$/BRL exchange rate of 3.60.

SRK undertook a sensitivity analysis of the PFS economic results. The Project’s economics are most sensitive to revenue and least sensitive to capital expenditure.

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26 Recommendations Based on the technical studies presented and the positive project economics SRK considers that the development of the Project from PFS to a FS level is warranted.

Further works include the following:

• It is recommended that a further round of resource drilling be undertaken to convert material currently classified as Inferred Mineral Resource into Indicated Mineral Resource or a higher resource classification. • The assayed Geotechnical holes should be included in the next Mineral Resource update. • Further work should be under taken to define higher grade shoots in the orebody. These are likely to have a significant effect on the economics of the deposit. • With further drilling the resource model will need to be updated. This update should include the use of smaller blocks (5 m×5 m × 5 m) to provide better granularity of the resource. As part of an updated Mineral Resource, the constraining wireframes used should be reviewed and revised where appropriate. Improvements to the understanding of the oxidation zones in the deposit would be very useful as they are likely to affect processing costs and recoveries. SRK notes that the smaller block size will need to be geostatistically supported by the drill spacing. • If there is a reasonable spread of assay results for elements other than gold, these should, after appropriate validation, be modelled especially if links can be made between associated mineralization and processing cost and recoveries. Similarly, the density (specific gravity, SG) should be included in the model as an estimated value, rather than a series of assigned values based on oxidation zone. • Further work to define the nature and extent of backfill in the historical WMC pits will allow better definition of the waste/ resource boundary. Such work may entail a LiDAR survey and a bathymetric survey in the old pits. • It is further suggested that when the resource is remodelled, both inverse distance squared (IDS) and Ordinary kriging (ok) interpolations should be carried out and compared to the Multiple Indicator Kriging (MIK) method with the modelling process peer reviewed prior to publication. • For the PFS update, SRK has assumed that the footwall orebody plunge and foliation are coincident (the regional geological data suggests foliation dips in the 35°–55° range which would support this assumption); at the FS stage, it may be possible to refine this if more structural data is available. • There is ~30 m of oxide on the footwall side which will need to be checked in the FS design – it is expected that horizontal drains and berm drainage will be needed to prevent zones of erosion coalescing to create larger zones of failure and pit seepage which will result in higher water management costs. This design can be reassessed when accessing the old pits to identify suitable geometry. • Major structures need to be mapped in the old pit once access is re-established and used to develop a working structural geological model to assist pit design. • Standard ground control/ slope management procedures need to be adopted so that the design assumptions are validated during mining and the design is further optimized. Mapping of the footwall structures will be very important to maintain the optimal pit production as well as checking for the potential for adverse footwall structures which could be unstable. • Good quality blasting of final walls and major intermediate cutbacks will be critical to good performance; pre-splitting (or similar blasting techniques) should be adopted.

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• Amarillo is conducting ongoing Bond Work Index (BWI) testwork that will give guidance on the grind power demand. • The power demand for the nominal grind is a key element in the GTR model. Ongoing testwork is recommended. • SRK recommends that during any future Definitive Feasibility Study (DFS), a complete review of all operating and capital costs assumptions will be necessary to ensure a DFS level of estimation accuracy. All major equipment should be re-quoted and new contractor quotes obtained for any future DFS process. • A study to further refine loss and dilution assumptions in the Mineral Reserve model should be undertaken. SRK recommends the use of a Regularised Reserve Block Model in any future optimizations or feasibility studies. • Increasing the mining rate may improve the NPV. A better understanding of the potential for ramp congestion by increasing the size of the truck fleet, the availability of additional dig faces in the pit, along with capital implications should be developed.

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27 References

Australian Exploration Field Services Pty Ltd, 2017. Mineral Resource Update for Posse Deposit, report prepared for Amarillo Gold Corporation.

BVP Engenharia, 2011. Visita Tecnica e Avaliacao dos Dados Geotecnicos do Deposito Posse – Projeto Mara Rosa, report prepared for Amarillo Gold Corporation.

Caracle Creek International Consulting, 2008. Independent Technical Report and Preliminary Economic Assessment, Mara Rosa Gold Property, Goiás State, Brazil, report prepared for Amarillo Gold Corporation.

Coffey, 2011a. Pre-feasibility Study, Mara Rosa Project, report prepared for Amarillo Gold Corporation.

Coffey, 2011b. Geotechnical Report for Pre-feasibility Study. Mara Rosa Project, report prepared for Amarillo Mineracao do Brasil Ltda, 2011.

Coffey, 2011c. Mara Rosa Metallurgical Testwork Report. Mara Rosa Pre-Feasibility Study, report prepared for Amarillo Gold Corporation.

Hidrovia. Estudos Hidrogeológicos e Hidrológicos Básicos ‘PFS’ Mina De Posse, Mara / Go. Prepared by Hidrovia Hidrogeologia e Meio Ambiente Ltda for Coffey Mining Pty Ltd and Amarillo Gold Corporation, 2011.

Hoogvliet Contract Services & Australian Exploration Field Services Pty Ltd, 2011. Report on Independent Site Visit and Resource Estimate, Posse Deposit, report prepared for Amarillo Gold Corporation.

Onix Engenharia e Consultoria Ltda. 2011. Estudo de Pré-Viabilidade. Mara Rosa Project, report prepared for and in collaboration with Amarillo Gold Corporation.

SRK Consultores do Brasil Ltda, 2017a. Projeto Conceitual da Barragem de Disposição de Rejeitos – Mina de Posse – Projeto Mara Rosa.

SRK Consultores do Brasil Ltda, 2017b. Posse Mine project – 2017 PFS Update, report prepared for Amarillo Gold Corporation.

Whittle Consulting Pty Ltd, 2018. Enterprise Optimization for the Amarillo Posse Project, prepared by Whittle Consulting.

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PO Box 943 West Perth WA 6872, Australia

T: +61 8 9288 2000 F: +61 8 9288 2001 E: [email protected]

www.asia-pacific.srk.com

28 Certificate of Qualified Person

a) I, Anthony Stepcich, am a Principal Consultant (Project Evaluations) with SRK Consulting (Australasia) Pty Ltd with a business address at PO Box 943, West Perth, WA, 6872, Australia.

b) This certificate applies to the technical report entitled ‘Technical Update on the Posse Gold Project, Brazil, September 2018’, dated 12 September 2018 (the ‘Technical Report’).

c) I am a graduate of Ballarat University with a Bachelor’s degree in Mining Engineering,1992. I am a graduate of Curtin University with a Master of Science Degree in Mineral Economics, 1997. I am a graduate of FINSIA with a Graduate Diploma in Finance and Investment, 2002. I am a Fellow and Chartered Professional in good standing of the Australasian Institute of Mining and Metallurgy, FAusIMM(CP), Membership number #110954. My relevant experience is that of a Mining Engineer with 25 years’ experience in the mining industry. I have gained both underground and open pit metalliferous experience, and open pit coal experience. I specialise in open pit design and scheduling and project evaluations. I am a Competent Person for the reporting of Ore Reserves in accordance with JORC Code (2012). I am also a Specialist in accordance with the VALMIN Code (2015) for the public reporting of valuations across multiple commodities. I have experience working in Australia and Indonesia. I am a ‘Qualified Person’ for purposes of National Instrument 43-101 (the ‘Instrument’).

d) I have not personally yet undertaken a site visit to the Posse Project in Brazil. The site visit was deferred. I am currently planning to visit the Posse Project in September / October 2018. After the proposed site visit, this Report will be updated with any material changes and refiled with SEDAR.

e) I am responsible for all items of the Technical Report, with the exception of Section 14.

f) I am independent of Amarillo Gold Corporation as defined by Section 1.5 of the Instrument.

g) I have no prior involvement with the Property that is the subject of the Technical Report.

h) I have read the Instrument and the Technical Report, and the Technical Report has been prepared in compliance with the Instrument.

i) At the effective date of the technical report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Anthony Stepcich, FAusIMM(CP) Principal Consultant (Project Evaluations) SRK Consulting (Australasia) Pty Ltd Dated 12 September 2018

STEP/WHIT/wulr ARO001_Posse Gold Project - NI 43 101 Technical Report_Rev2.docx 12 September 2018 Level 1, 10 Richardson Street West Perth WA 6005, Australia

PO Box 943 West Perth WA 6872, Australia

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29 Certificate of Authorship I, Gregory Keith Whitehouse, MAusIMM, CP (Geo), as author of Section 14 of the Technical Report entitled ‘Technical Update on the Posse Gold Project, Brazil, September 2018’ prepared for Amarillo Gold Corporation. (‘Issuer’), do hereby certify that:

I am currently employed as a Geological Consultant and Director of Australian Exploration Field Services Pty Ltd. with offices at 14 Brodie Street, Bendigo, Victoria, Australia 3550.

1) I graduated with a Bachelor of Science degree in Geology from Victoria University, Wellington, New Zealand in 1975.

2) I am a Professional Geoscientist enrolled with the Australasian Institute of Mining and Metallurgy (Member # 107612) for over 27 years.

3) I have worked as a geologist for a total of 38 years since my graduation from university. My experience has covered exploration and technical management work on projects in South East Asia, Africa, South America and Australia including 14 years as an independent geology consultant. I have worked on copper, gold, base metals, tin, iron and specialist metals and have authored or contributed as a ‘qualified person’ to seven NI 43 101 reports and numerous reports under the JORC Code.

4) I have read the definition of ‘qualified person’ set out in National Instrument 43-101 (NI 43-101) and certify that, by reason of my education, affiliation with a professional association as defined in NI 43-101, and past relevant work experience, I fulfil the requirements to be a ‘Qualified Person’ for the purposes of NI 43-101.

5) I am responsible for entirety of Section 14 of the Technical report entitled ‘Technical Update on the Posse Gold Project, Brazil, September’ dated 12 September 2018.

6) I visited the Mara Rosa property in July 2012.

7) As of the effective date of the certificate, to the best of my knowledge, information, and belief, the sections of the Technical Report that I have authored contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

8) I have no personal knowledge, as of the date of this certificate, of any material fact or material change which is not reflected in this Technical Report.

9) I am independent of the Issuer, Amarillo Gold Corporation, applying all the tests in Section 1.5 of the NI 43-101 instrument.

10) I have acted as an Independent expert in the preparation of a number of Resource Reports over the Mara Rosa property.

11) I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

Mr Keith Whitehouse, MAusIMM CP(Geo) Director Australian Field Exploration Services Pty Ltd (AEFS) Dated 12 September 2018

STEP/WHIT/wulr ARO001_Posse Gold Project - NI 43 101 Technical Report_Rev2.docx 12 September 2018