NI 43-101 TECHNICAL REPORT
ALTAN TSAGAAN OVOO GOLD PROJECT
Tsagaan Ovoo, Dornod, Mongolia
Prepared for:
Steppe Gold LLC Shangri-La office 1201, Shangri-La office, Olympic street 19A, Sukhbaatar district-1, Ulaanbaatar Mongolia
Prepared by: Qualified Person: Oyungerel Bayanjargal, MAusIMM (CP) GSTATS Consulting LLC
and
Reviewed by: Qualified Persons:
Leonid Tokar, P.Eng, GSTATS Consulting LLC Tim Fletcher, P.Eng, DRA Americas Inc. David Frost, FAusIMM, B.Met Eng., DRA Americas Inc. Dr. Martin Stapinsky, P.Geo, DRA Americas Inc.
Effective date: September 6, 2017 Issue date: October 4, 2017 SIGNATURE PAGE
To accompany the report entitled NI 43-101 TECHNICAL REPORT ALTAN TSAGAAN OVOO GOLD PROJECT, Tsagaan Ovoo, Dornod, Mongolia dated October 4, 2017.
Qualified Person
Date: October 4, 2017 Oyungerel Bayanjargal, MAusIMM (CP) Principal Consultant GSTATS Consulting LLC CERTIFICATE OF QUALIFIED PERSON
To accompany the report entitled NI 43-101 TECHNICAL REPORT ALTAN TSAGAAN OVOO GOLD PROJECT, Tsagaan Ovoo, Dornod, Mongolia dated October 4, 2017.
I, Oyungerel Bayanjargal, do hereby certify that:
1. I am Director and Principal Consultant for GSTATS Consulting LLC with an office at 407a, Peace Building, Sukhbaatar Street, Chingeltei, Ulaanbaatar, Mongolia. 2. I hold the following academic qualifications: B.Sc in Mathematics, University of Education, Mongolia; B.Sc (Honours) in Geology, University of Science and Technology, Mongolia; M.Sc in Mineral Resource Exploration, ITC Twente University, The Netherlands. 3. I am a Member (No.301956) of the Australasian Institute of Mining and Metallurgy and accredited as a Chartered Professional under the discipline of Geology. 4. I have practiced my profession continuously in exploration and mining industry since my graduation from university. I have 24 years of experience in mineral exploration and mining. Since 2004, I’ve carried out Mineral Resource and Mineral Reserve estimates, and Pre- Feasibility and Feasibility Studies predominately in a gold mine as Senior mine geologist and Chief Mine geologist. My consulting experience since 2011 was mostly focused on gold, copper and base metal deposits. 5. I have read the definition of “qualified person” set out in the National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association and past relevant work experience, I fulfil the requirements to be a qualified person for the purposes of NI 43-101. 6. I am the author of this technical report and responsible for the preparation of Section 1 to Section 27. 7. I conducted a site inspection of the project on 23 August and 2 October of 2017. 8. I am independent of the issuer, Steppe Gold LLC. 9. I have not had any prior involvement with ATO project. 10. I have read NI 43-101 and Form 43-101F1 and the sections of the Technical Report for which I was responsible have been prepared in compliance with that instrument and form and 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.
Dated in Ulaanbaatar, Mongolia, this October 4, 2017.
Oyungerel Bayanjargal, MAusIMM (CP) Principal Consultant, GSTATS Consulting LLC CERTIFICATE OF QUALIFIED PERSON
To accompany the report entitled NI 43-101 TECHNICAL REPORT ALTAN TSAGAAN OVOO GOLD PROJECT, Tsagaan Ovoo, Dornod, Mongolia dated October 4, 2017.
I, Lenny Tokar, do hereby certify that:
1. I am an Associated Consultant for GSTATS Consulting LLC, 10 Old Oak Dr. #205, Buffalo Grove, IL, 60089, USA. 2. I hold the following academic qualifications: Master's Degree in Mining Engineering, University, Krivoy Rog, “Strategic Risk Management and Applied Optimization in Mine Design” McGill University, Montreal, Canada. 3. I am a Professional Mining Engineer of Ontario (No. 100014617). 4. I have practiced Mining Engineering since my graduation from university. I have more than 30 years of Mine Engineering experience. Since 2005, I’ve carried out Mineral Resource and Mineral Reserve estimates, Pre-Feasibility and Feasibility Studies predominately for the gold mining sector as Senior Mine Engineer. My consulting experience since 2015 was focused on gold, copper and base metal deposits. 5. I have read the definition of “qualified person” set out in the National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association and past relevant work experience, I fulfil the requirements to be a qualified person for the purposes of NI 43-101. 6. I am responsible for the review of the Sections 15, 16 and 22. 7. I am independent of the issuer, Steppe Gold LLC. 8. I have not had any prior involvement with ATO project. 9. I have read NI 43-101 and Form 43-101F1 and the sections of the Technical Report for review of which I was responsible have been prepared in compliance with that instrument and form and 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.
This 22nd day of January 22, 2018.
Leonid (Lenny) Tokar, P.Eng. Associated Consultant, GSTATS Consulting LLC
□RA
th This 26 day of February 2018.
Tim Fletcher, P. Eng. Project Manager DRA Americas Inc.
□RA 11) Neither I, nor any affiliated entity of mine, own, directly or indirectly, nor expect to receive, any interest in the properties or securities of Steppe Gold LLC, or any associated or affiliated companies; 12) Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three (3) years from Steppe Gold LLC, or any associated or affiliated companies; 13) I have read NI 43-101 and Form 43-l 0lFl and have reviewed the relevant sections involved in (as listed in point 6) in the Technical Report in compliance with NI 43- 101 and Form 43-1 0lFl; these Sections were prepared in conformitywith generally accepted - mining industry best practice, and as of the date of the certificate, to the best of my knowledge, information and belief, the Technical Report contains the scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
This 26th day of February 2018,
David Frost, F AusIMM, B. Met Eng. Vice President - Process Engineering DRA Americas Inc.
Page 2 of2 A日時A
CERTIFICATE OF QUALIFIED PERSON
To accompany the Report titled “NI 43" 1 0 1 TECHNICAL REPORT ALTAN TSAGAAN OVOO GOLD PROJECT, Tsagaan Ovoo’Domod, Mongolia” which is e節ective as of
September 6血, 2017 and issued on October 4th, 2017 (血e “Tec血ical Report”) prepared for
Steppe Gold LLC (the “Company’’)・
I, Martin Stapinsky, P. Geo, M.Sc., Ph.D., do hereby certify that: l)喜諾霊誤露l豊悪霊鵠講説霊窪霊 2)嘉宝器誓認諾認諾慧智慧豊霊霊露盤
a ph. D. in Earth Sciences (hydrogeoIogy) from the University of Ottawa in 2001 ; 3) k器霊宝霊諾‡盗,駕欝箸bec (#794) and the
4) I have worked as a HydrogeoIogist and in Enviroment continuously since my 5)豊富羅霊豊諾記譜岩盤豊霊諾graduation from university;
association and past relevant wock experience, I細則the requlrementS tO be an independent qua雌ed person for the puaposes ofNI 43-101 ; 6)岩盤認諾霊黙諾藍慧嵩豊隷書藍霊
the enviromental studies, Pem舶ng, SOCial and cormu血ty aspects ofthe prQject;
7) I have not visited the Altan Tsagaan Ovoo PrQject (ATO) site;
8) I have no personal knowledge as ofthe date of址s cert:胞cate ofany material fact or change, Which is not reflected in this Report; 9)慧霊土器嵩認諾霊詫霊諾。豊諾諾
entity or empIoyee of Steppe Gold LLC, Or any aSSOCiated or a珊iated entities; 10)霊器詳警謹認諾筈霊誤認詰ま藍蒜嵩
associated or a節1iated companies; 11)謹言葦品諾露盤詮議諸富諒等霊諾
a触Iiated companies;
P呼l of2 □RA 12) I have read NI 43-101 and Form 43-lOlFl and have reviewed the relevant sections involved in (as listed in point 6) in the Technical Report in compliance with NI 43- 101 and Form 43-1 OlFl; these Sections were prepared in conformity with generally accepted mining industry practice, and as of the date of the certificate, to the best of my knowledge, information and belief, the Technical Report contains the scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
This 26th day of February 2018,
Martin Stapinsky, P.Geo. M.Sc., Ph.D. Senior ProfessionalGeologist Met-Chem, a division of DRA Americas Inc.
Page 2 of2 NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT Table of Contents
1. SUMMARY ...... 11 1.1 INTRODUCTION ...... 11 1.2 PROPERTY DESCRIPTION AND OWNERSHIP ...... 11 1.3 GEOLOGY AND MINERALIZATION ...... 11 1.4 EXPLORATION AND DRILLING ...... 11 1.5 DATA VERIFICATION ...... 12 1.6 METALLURGICAL TESTS AND PROCESSING ...... 12 1.7 MINING METHODS AND PIT OPTIMIZATION ...... 12 1.8 RECOVERY METHODS ...... 13 1.9 MINE PRODUCTION SCHEDULE ...... 13 1.10 CAPITAL COSTS ...... 13 1.11 OPERATING COSTS ...... 14 1.12 REVENUE ...... 14 1.13 MINERAL RESERVES AND MINERAL RESOURCES SUMMARY ...... 15 1.14 CONCLUSIONS AND RECOMMENDATIONS ...... 15 2 INTRODUCTION...... 18 2.1 ISSUER FOR WHOM THIS REPORT WAS PREPARED ...... 18 2.2 TERMS OF REFERENCE AND PURPOSE OF REPORT ...... 18 2.3 UNITS OF MEASURE AND CURRENCY...... 18 2.4 SOURCES OF INFORMATION AND STUDY PARTICIPANTS ...... 18 2.5 PERSONAL SITE INSPECTIONS ...... 19 2.6 ABBREVIATIONS AND NOMENCLATURE ...... 20 3 RELIANCE ON OTHER EXPERTS ...... 21 4 PROPERTY DESCRIPTION AND LOCATION ...... 22 4.1 PROPERTY OWNERSHIP AND BOUNDARIES ...... 22 4.2 INVESTMENT AGREEMENT (IA) ...... 23 4.3 ROYALTIES ...... 23 4.4 REQUIRED PERMITS ...... 24 4.5 SOCIO-ECONOMIC IMPACTS ...... 26 4.6 COMMUNITY POLICY, PLANNING, CONSULTATION AND ENGAGEMENT ...... 26 5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY ...... 27 5.1 TOPOGRAPHY, ELEVATION, VEGETATION AND WILD LIFE ...... 27 5.2 PROPERTY ACCESS ...... 27 5.3 REGIONAL POPULATION AND INFRASTRUCTURE ...... 27 5.4 CLIMATE AND OPERATING SEASON ...... 27 5.5 WATER ...... 28 5.6 LAND USE ...... 28 5.7 RISK ASSESSMENTS ...... 28 5.8 CLOSURE AND RECLAMATION ...... 28 6 HISTORY ...... 29 6.1 HISTORY AND LAND HOLDINGS ...... 29 6.2 PREVIOUS MINERAL RESERVES AND MINERAL RESOURCES ...... 30 7 GEOLOGIAL SETTING AND MINERALIZATION ...... 32 7.1 REGIONAL GEOLOGICAL SETTING ...... 32 7.2 DISTRICT GEOLOGY AND MAGMATISM ...... 33 7.2.1 Mineralization in the District Surrounding ATO ...... 36 7.3 GEOLOGY OF THE ATO DEPOSIT ...... 36 Page 1 of 209
NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT
7.3.1 Geology ...... 36 7.3.2 Controlling Factors ...... 37 7.3.3 Silicate Alteration in Mineralized Pipes ...... 38 7.3.4 Mineralization at ATO ...... 45 8 DEPOSIT TYPE ...... 48 8.1 GEOLOGICAL MODEL FOR ATO DISTRICT ...... 48 8.2 GENESIS OF THE ATO DEPOSIT ...... 50 8.2.1 Formation of Breccia Pipe ...... 50 8.3 GEOCHEMICAL ZONATION ...... 51 8.4 DEPOSIT TYPE ...... 51 9 EXPLORATION ...... 53 9.1 PREPARATION STAGE ...... 53 9.2 FIELD WORK ...... 53 9.2.1 Geological Mapping and Prospecting Traverses ...... 53 9.2.2 Field Sampling for Geochemical Analysis ...... 53 9.2.3 Other Sampling tasks ...... 55 9.3 EXPLORATION RESULTS ...... 59 9.3.1 Geochemical Analysis ...... 59 9.3.2 Trenching ...... 65 9.3.3 Geophysical Survey ...... 66 9.3.4 Groundwater Water Exploration Results ...... 73 10 DRILLING ...... 74 10.1 DRILL PROGRAMS ...... 74 10.2 DRILL ORIENTATIONS ...... 74 10.3 DRILL CONTRACTORS ...... 75 10.4 CORE TRANSPORT AND STORAGE ...... 76 10.5 GEOLOGICAL LOGGING ...... 76 10.6 CORE RECOVERIES ...... 76 10.7 COLLAR AND DOWNHOLE SURVEYS ...... 76 10.8 CORE SAMPLING ...... 77 11 SAMPLE PREPERATION, ANALYSES AND SECURITY...... 78 11.1 SAMPLE PREPARATION ...... 78 11.2 LABORATORY ANALYSES ...... 79 11.3 QUALITY ASSURANCE AND QUALITY CONTROL (QAQC) ...... 79 11.3.1 Blank Samples ...... 80 11.3.2 Standard Samples...... 80 11.3.3 Duplicate Samples ...... 89 11.3.4 QAQC for Laboratories ...... 94 11.4 SAMPLE SECURITY ...... 98 11.5 CONSIDERATIONS AND CONCLUSIONS ...... 98 12 DATA VERIFICATION ...... 99 12.1 GENERAL ...... 99 12.2 DRILL HOLE DATABASE VALIDATION ...... 99 12.3 ASSESSMENT OF DATABASE ...... 100 13 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 101 13.1 XSTRATA PROCESS SUPPORT (XPS) METALLURGICAL OXIDE TESTING ...... 101 13.1.1 Mineralogical results ...... 101 13.1.2 Metallurgical Testwork Results ...... 104 13.1.3 Gravity Recoverable Gold ...... 104 13.1.4 Bottle Roll Leach Tests ...... 105 13.2 G&T METALLURGICAL OXIDE TESTING PROGRAM ...... 106
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT
13.2.1 Sampling ...... 107 13.2.2 Ore Characteristics ...... 110 13.2.3 Mineral Distributions ...... 111 13.2.4 Recovery Response to Grind Size – Master Composite 1 ...... 111 13.3 HEAP LEACH AMENABILITY TEST WORK – COLUMN LEACH COMPOSITE ...... 112 13.3.1 Natural Percolation Rate Test Work...... 112 13.3.2 Column Cyanidation Leach Test Work ...... 113 13.4 LARGE-SCALE COLUMN TESTS ...... 113 13.5 CONCLUSIONS ...... 117 14 MINERAL RESOURCE ESTIMATES ...... 118 14.1 DATABASES ...... 118 14.1.1 Drill Hole Database ...... 118 14.1.2 Bulk Density Data ...... 118 14.2 GEOLOGY AND GRADE SHELL MODELLING ...... 119 14.2.1 Lithology Modelling ...... 119 14.2.2 Fault Modelling ...... 120 14.2.3 Grade Shells ...... 120 14.3 BLOCK MODEL ...... 120 14.3.1 Block Model Properties ...... 120 14.3.2 Block Model Variables Definition and Tagging ...... 121 14.4 DEPOSIT STATISTICS ...... 122 14.4.1 Compositing ...... 122 14.4.2 Exploratory Data Analysis ...... 124 14.4.3 Grade Capping and Evaluation of Extreme Grades ...... 132 14.4.4 Metals Correlation ...... 134 14.5 GEOSPATIAL ANALYSIS ...... 134 14.6 GRADE BLOCK MODELS ...... 136 14.7 BLOCK MODEL VALIDATION ...... 138 14.7.1 Visual validation ...... 138 14.7.2 Grade Swath Plots and Local Bias Checks ...... 138 14.7.3 Global Mean Statistics ...... 140 14.8 GOLD EQUIVALENT CALCULATION ...... 140 14.9 MINERAL RESOURCE CLASSIFICATION...... 141 14.9.1 Mineral Resource Classification Model ...... 141 14.9.2 Mineral Resource Classification, Category and Definition ...... 142 14.9.3 Estimation ...... 142 14.9.4 Derivation of Cut-off Grades ...... 145 14.9.5 Mineral Resource Statement ...... 146 14.10 FACTORS THAT COULD AFFECT THE MINERAL RESOURCE ESTIMATE ...... 147 15 MINERAL RESERVE ESTIMATES...... 148 15.1 PIT OPTIMIZATION ...... 148 15.1.1 Economic Parameters...... 148 15.1.2 Cut-off grades...... 149 15.1.3 Slope Parameters ...... 149 15.1.4 Pit Optimization Result ...... 149 15.1.5 Ultimate Pit Limit Selection ...... 151 15.2 PIT DESIGN ...... 151 15.2.1 Bench Height ...... 151 15.2.2 Pit design Slopes ...... 151 15.2.3 Haulage Roads ...... 152 15.2.4 Ultimate Pit ...... 152 15.3 DILUTION AND LOSS ...... 153 15.4 MINERAL RESERVES ESTIMATE ...... 153 Page 3 of 209
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15.5 MINERAL RESERVES STATEMENT ...... 154 15.5.1 In-pit Inferred Resources ...... 155 16 MINING METHODS ...... 156 16.1 MATERIAL TYPES ...... 156 16.1.1 Waste Material ...... 156 16.1.2 Ore Material types ...... 156 16.2 MINING METHOD ...... 156 16.3 MINE-WASTE FACILITIES ...... 156 16.4 MINE PRODUCTION SCHEDULE ...... 156 16.5 MINING EQUIPMENT ...... 158 16.6 MINING OPERATIONS ...... 158 16.6.1 Drilling and Blasting ...... 158 16.6.2 Loading and Hauling ...... 158 16.6.3 Mine Personnel ...... 159 17 RECOVERY METHODS ...... 160 17.1 PROCESS DESIGN CRITERIA ...... 160 17.2 PRODUCTION RATE AND PRODUCTS ...... 162 17.3 PROCESS DESCRIPTION ...... 162 17.3.1 Crushing ...... 163 17.3.2 Heap Leach Facility ...... 163 17.3.3 ADR Gold Recovery Plant ...... 165 18 PROJECT INFRASTRUCTURE ...... 168 18.1 FACILITIES...... 168 18.1.1 Processing Plant ...... 168 18.1.2 Laboratory ...... 168 18.1.3 Warehouse and maintenance areas ...... 169 18.1.4 Office building ...... 170 18.1.5 Fuel depot ...... 171 18.1.6 Explosive storage ...... 171 18.1.7 Camp ...... 171 18.2 WATER SUPPLY ...... 172 18.3 POWER SUPPLY ...... 172 18.4 ROADS ...... 173 18.5 COMMUNICATIONS ...... 173 19 MARKET STUDIES ...... 174 20 ENVIROMENTAL STUDIES, PERMITTING, AND SOCIAL IMPACT ...... 175 20.1 MONITORING SYSTEM ON ENVIRONMENTAL PROTECTION ...... 175 20.2 LIST OF MINING LICENSES AND REGULATIONS TO BE APPLIED ...... 176 20.3 ENVIRONMENTAL IMPACT: IDENTIFYING, REDUCING AND ERADICATING PLAN ...... 177 20.3.1 Soil layers ...... 177 20.3.2 Plants ...... 178 20.3.3 Water ...... 179 20.3.4 Air quality ...... 179 20.3.5 Fauna ...... 179 20.3.6 Archeological findings ...... 179 20.3.7 Public collaboration ...... 180 20.3.8 Other environmental issues ...... 180 20.4 GENERAL APPROACH TO MANAGEMENT OF ACID ROCK DRAINAGE ...... 180 20.4.1 Waste Rock Management Plan ...... 181 20.5 NEUTRALIZATION AND STORAGE OF CHEMICAL SOLUTION AND POISONOUS SUBSTANCES ...... 181 20.5.1 Natrium cyanide ...... 181
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20.5.2 Hydrochloric acid ...... 182 20.5.3 Copper sulfate ...... 182 20.6 PREVENTING POISONOUS SUBSTANCE FROM SEEPING INTO SOIL ...... 182 20.7 MEASURES TO TAKE DURING SPECIAL CONDITIONS ...... 184 20.8 MINING CLOSURE, REHABILITATION, COST ESTIMATION...... 184 20.8.1 First phase activities of mining closure ...... 184 20.8.2 Intense rehabilitation activity ...... 184 20.8.3 Rehabilitation activity ...... 186 21 CAPITAL AND OPERATING COSTS ...... 188 21.1 CAPITAL COSTS (INVESTMENT) ...... 188 21.2 OPERATING COSTS (OPERATIONAL EXPENSES) ...... 189 22 ECONOMIC ANALYSIS ...... 195 22.1 REVENUE ...... 195 22.2 CASH FLOW ANALYSIS ...... 196 22.3 INCOME STATEMENT AND PROFIT RATIOS ...... 198 22.4 DISCOUNTED CASH FLOW (AFTER TAX) ...... 198 22.5 STREAM ANALYSIS ...... 200 23 ADJACENT PROPERTIES ...... 204 24 OTHER RELEVANT DATA AND INFORMATION ...... 205 25 INTERPRETATION AND CONCLUSIONS ...... 206 25.1 CONCLUSIONS ...... 206 25.2 RISK ASSESSMENT ...... 207 26 RECOMMENDATIONS ...... 208 27 REFERENCES ...... 209
Page 5 of 209
NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT List of Figures
Figure 4.1: Location map for the ATO project ...... 22 Figure 6.1: ATO Mineral Resources in 2016 AIF by Centerra Gold Inc. SEDAR ...... 30 Figure 7.1: Location of the Mongol-Okhotsk belt and Onon precious-base-metal province residing between the Siberian and North China Plates (CGM 2012 ER)...... 32 Figure 7.2: District geological map of ATO project ...... 35 Figure 7.3: Mineralized pipes at ATO deposit (View to WNW) ...... 36 Figure 7.4: Plan map of ATO mineralized pipes...... 36 Figure 7.5: Longitudinal cross section through Pipe 1 and Pipe 2. Au and Ag grades in right and left sides respectively...... 37 Figure 7.6: Schematic cross section through Pipe 1 at ATO (A) showing relative abundance of alteration minerals with depth (B) ... 38 Figure 7.7: Back scattered electron micrographs in polished thin section...... 39 Figure 7.8: Aspects of various alteration assemblages typically present in drill core at ATO ...... 39 Figure 7.9: Photographs of outcrops of silica cap rock at Pipe 1 at ATO...... 41 Figure 7.10: Relict reticulated mats of silicified reeds siliceous sinter. Plane polarized transmitted light...... 42 Figure 7.11: Photomicrographs in reflected light of particles of free gold (Au) in surface grab samples ...... 42 Figure 7.12: Photographs of silica cap rock in oxidized drill core from Pipe 1 ...... 43 Figure 7.13: General fabric of mineralized breccia in drill holes from Pipe 2, ATO...... 44 Figure 7.14: General styles of mineralized rock at ATO ...... 44 Figure 7.15: Sulfur fugacity versus temperature diagram ...... 47 Figure 8.1: Cartoon model of the Jurassic intrusives in the ATO district...... 49 Figure 8.2: Model of formation of ATO deposit ...... 50 Figure 8.3: Section Model of ATO deposit indicating relationship of the deposit with a copper porphyry intrusion ...... 51 Figure 8.4: Nature and types of epithermal mineralization ...... 52 Figure 9.1: Map showing soil sample locations on gold patterns in ATO District during 2010-2014 (CGM 2014ER)...... 55 Figure 9.2: Combined dispersion halo at the ATO deposit ...... 60 Figure 9.3: Gold and other elements in Pipe 1 and their correlation at depth ...... 62 Figure 9.4: Gold and other elements in Pipe 2 and their correlation at depth ...... 63 Figure 9.5: Gold and other elements in Pipe 4 and their correlation at depth...... 65 Figure 9.6: Magnetic survey of surrounding area of ATO deposit ...... 67 Figure 9.7: Detailed magnetic survey of ATO deposit, Au distribution presented...... 68 Figure 9.8: D-D IP, 100–m chargeability. Dashed blue lines, lines of IP profiles...... 68 Figure 9.9: D-D IP, 100–m resistivity. Dashed blue lines, lines of IP profiles...... 69 Figure 9.10: D-D IP Resistivity map at the ATO deposit. Au distribution presented...... 69 Figure 9.11: D-D IP Chargeability map at the ATO deposit. Au distribution presented...... 70 Figure 9.12: A cross section for D-D IP Resistivity map at the ATO deposit. Au distribution presented...... 70 Figure 9.13: A cross section for D-D IP Chargeability map at the ATO deposit. Au distribution presented...... 70 Figure 9.14: A cross section for D-D IP Chargeability map at the ATO deposit (Pipe 1, 2, and 4). Au distribution presented...... 71 Figure 9.15: Gravity survey of surrounding area of ATO deposit ...... 72 Figure 9.16: Gravity map at the ATO deposit. Au distribution presented...... 72 Figure 10.1: Location map of drill holes ...... 75 Figure 11.1: Sample preparation procedure ...... 78 Figure 11.2: Assay results of gold blank samples ...... 80 Figure 11.3: Assay results of standard samples from Ore Research & Exploration ...... 81 Figure 11.4: Assay results of low-gold grade standard samples of Geostats ...... 82 Figure 11.5: Rocklab assay results of standard samples on low-grade oxidized mineralization ...... 83 Figure 11.6: Rocklab assay results of standard samples on medium-grade oxidized mineralization ...... 84 Figure 11.7: Rocklab assay results of standard samples on medium-grade fresh mineralization ...... 84 Figure 11.8: Rocklab assay results of standard samples on high-grade fresh mineralization ...... 84 Figure 11.9: Blank standard results for Ag ...... 85 Figure 11.10: Low-grade standard sample results for Ag ...... 85 Figure 11.11: High-grade standard sample results for Ag ...... 86 Figure 11.12: Blank standard results for Cu ...... 86 Figure 11.13: Low-grade standard sample results for Cu ...... 86 Figure 11.14: High-grade standard sample results for Cu ...... 87 Figure 11.15: Blank standard results for Pb ...... 87 Figure 11.16: Low-grade standard sample results for Pb ...... 87 Figure 11.17: High-grade standard sample results for Pb ...... 88 Figure 11.18: Blank standard results for Zn...... 88 Page 6 of 209
NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT
Figure 11.19: Low-grade standard sample results for Zn ...... 88 Figure 11.20: High-grade standard sample results for Zn ...... 89 Figure 11.21: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 20 ppm Au ...... 90 Figure 11.22: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 20 ppm Au ...... 91 Figure 11.23: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 80 ppm Ag ...... 91 Figure 11.24: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 80 ppm Ag ...... 92 Figure 11.25: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 6,000 ppm Cu .. 92 Figure 11.26: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 6,000 ppm Cu ...... 92 Figure 11.27: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 30,000 ppm Pb 93 Figure 11.28: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 30,000 ppm Pb ...... 93 Figure 11.29: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 70,000 ppm Zn 94 Figure 11.30: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 70,000 ppm Zn ...... 94 Figure 11.31: Results of gold assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS ...... 96 Figure 11.32: Results of silver assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS ...... 96 Figure 11.33: Results of copper assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS ...... 97 Figure 11.34: Results of lead assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS ...... 97 Figure 11.35: Results of zinc assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS ...... 98 Figure 13.1: Modal Mineralogy of Oxide, Upper Transition and Lower Sulphide Zones ...... 102 Figure 13.2: Mineral distribution by class of association Master composite 1, (78μm k80)...... 109 Figure 13.3: Mineral Distribution by Class of Association ...... 111 Figure 13.4: Flowsheet - Master composite 1 ...... 111 Figure 13.5: Rate of extraction ...... 112 Figure 13.6: Large scale column test at Boroo Mine ...... 114 Figure 13.7: Large scale column test flowsheet ...... 114 Figure 13.8: Large scale column testing process ...... 115 Figure 13.9: Gold recovery curve (Testing for 12.5mm) ...... 116 Figure 13.10: Gold recovery curve (Testing for 25mm) ...... 117 Figure 13.11: Gold recovery curve (Testing for 50mm) ...... 117 Figure 14.1: Histogram of Bulk Density of ATO. A: All; B: Oxidized; C: Transition; and D: Fresh samples ...... 119 Figure 14.2: Diorite Dyke Model and Grade Shells ...... 120 Figure 14.3: Fault Model and Grade shells ...... 120 Figure 14.4: Block model dimension in GEMS ...... 121 Figure 14.5: Rock type block model with drill hole oxide information (Cross section 9SE) ...... 122 Figure 14.6: Histogram of Sample Length at ATO deposit ...... 123 Figure 14.7: Au (ppm) Histogram for Pipe 1, Pipe 2 and Pipe 4 ...... 126 Figure 14.8: Ag (ppm) Histogram for Pipe 1, Pipe 2 and Pipe 4 ...... 127 Figure 14.9: Pb (ppm) Histogram for Pipe 1, Pipe 2 and Pipe 4 ...... 128 Figure 14.10: Zn (ppm) Histogram for Pipe 1, Pipe 2 and Pipe 4 ...... 129 Figure 14.11: Box and Whisker Plot for Au ...... 130 Figure 14.12: Box and Whisker Plot for Ag ...... 130 Figure 14.13: Box and Whisker Plot for Pb ...... 131 Figure 14.14: Box and Whisker Plot for Zn ...... 131 Figure 14.15: Au Probability Plot and Mean-Variance Plot for Pipe 1 ...... 132 Figure 14.16: Au Probability Plot and Mean-Variance Plot for Pipe 2 ...... 132 Figure 14.17: Au Probability Plot and Mean-Variance Plot for Pipe 4 ...... 133 Figure 14.18: Variogram Models for Au in pipes ...... 135 Figure 14.19: Variogram Models for Ag in pipes ...... 135 Figure 14.20: Variogram Models for Pb in pipes ...... 136 Figure 14.21: Variogram Models for Zn in pipes ...... 136 Figure 14.22: A Cross Section view showing estimated block grades of the ATO Pipes...... 137 Figure 14.23: 3D view of estimated Au block grades of the ATO Pipes ...... 137 Figure 14.24: Cross section for visual validation on Section 9SE ...... 138 Figure 14.25: Swath Plots for Au, Ag, Pb and Zn for the deposit: Red-composites, blue-OK Model, green-ID Model ...... 139 Figure 14.26: Mineral Resource classification block model, Pipe 1: Red – Measured; Blue – Indicated; Green – Inferred resources. 141 Figure 15.1: Graph of Whittle Results (AuEq cut-off 0.3 g/t) ...... 150 Page 7 of 209
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Figure 15.2: Ultimate Whittle pit shell and pipes ...... 151 Figure 15.3: Ultimate pit design (Surpac mine design) ...... 152 Figure 15.4: Ultimate pit design ...... 153 Figure 16.1: Mine production plan ...... 157 Figure 16.2: Gold and Silver in ROM ore...... 157 Figure 17.1: Diagram for the process flows ...... 162 Figure 17.2: Schematic flowsheet for the process flows ...... 163 Figure 17.3: HLF plan ...... 164 Figure 17.4: HLF Cross section ...... 165 Figure 18.1: Processing Plant ...... 168 Figure 18.2: Laboratory building ...... 169 Figure 18.3: Warehouse and maintenance area ...... 170 Figure 18.4: Office building ...... 170 Figure 18.5: Office plan ...... 171 Figure 18.6: Water supply plan ...... 172 Figure 19.1: LME historical Gold price chart ...... 174 Figure 19.2: LME historical Silver price chart ...... 174 Figure 20.1: Damaged land from mine activities ...... 184 Figure 21.1: OPEX per ounce, USD...... 193 Figure 21.2: OPEX per ton of ore, USD ...... 193 Figure 22.1: Gold produced, oz...... 195 Figure 22.2: Silver produced, oz...... 195 Figure 22.3: Accumulated Cash flow (non-discounted), thousand USD...... 196 Figure 22.4: Discounted cash flow (after tax), thousand USD ...... 199 Figure 22.5: Sensitivity analysis of IRR ...... 200 Figure 22.6: Discounted cash flow with stream, thousand USD ...... 202
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT List of Tables
Table 1.1: Mine Production schedule ...... 13 Table 1.2: Project investment summary ...... 14 Table 1.3: Project operational cost and general expenses ...... 14 Table 1.4: Revenue ...... 14 Table 1.5: ATO Deposit Mineral Reserve and Mineral Resource Summary (at August 21, 2017) ...... 15 Table 2.1: List of Abbreviations ...... 20 Table 4.1: ATO Project license boundary coordinates ...... 23 Table 4.2: Mongolian Government State sliding scale royalty for Au and Ag ...... 24 Table 4.3: Mongolian Government State sliding scale royalty for Pb and Zn ...... 24 Table 9.1: Schedule of metallurgical samples ...... 56 Table 9.2: Sets of samples for initial metallurgical test ...... 57 Table 9.3: Sets of samples for second metallurgical test ...... 58 Table 9.4: Correlations of elements in geochemical dispersion halo of Pipe 1 ...... 61 Table 9.5: Correlations of elements in geochemical dispersion halo of Pipe 2 ...... 61 Table 9.6: Correlations of elements in geochemical dispersion halo of Pipe 4 ...... 61 Table 9.7: Correlations of elements in core samples from Pipe 1 ...... 62 Table 9.8: Correlations of elements in core samples from Pipe 2 ...... 63 Table 9.9: Correlations of elements in core samples from Pipe 4 ...... 64 Table 9.10: Significant trenching results ...... 66 Table 10.1: Exploration Drill Hole Summary Table ...... 74 Table 11.1: Number of samples for quality assurance and quality control of the ATO project ...... 80 Table 11.2: Standard samples from the laboratory of Ore Research and Exploration ...... 81 Table 11.3: Standard samples from the laboratory of Geostats ...... 82 Table 11.4: Rocklab Standard samples ...... 83 Table 11.5: Standard samples of polymetallic elements obtained from Geostats Pty ...... 85 Table 11.6: Comparison of primary and duplicate samples ...... 89 Table 11.7: Comparison of variations of primary and duplicates samples assay results ...... 90 Table 11.8: Changes in correlation of laboratories ACTLABS and SGS ...... 95 Table 11.9: Comparison of assay results of ACTLABS and SGS ...... 95 Table 13.1: Modal Mineralogy of Oxide, Upper Transition and Lower Sulphide Zones ...... 102 Table 13.2: Average Grain Size Measurements in Polished Thin Sections and 6 Mesh Material ...... 103 Table 13.3: QEMSCAN analysis result on Oxide Zone ...... 103 Table 13.4: Oxide (ATO-01) Composite ...... 104 Table 13.5: Minor Elements Analysis for Composites ...... 104 Table 13.6: DOE for Leaching Tests on Upper Oxide (ATO-01) ...... 105 Table 13.7: Results of the DOE at 48 hours of leaching ...... 106 Table 13.8: Master Composite 1 ...... 107 Table 13.9: Column Leach Composites ...... 108 Table 13.10: Mineral Composition and Elemental Deportment of the Oxide Composite ...... 109 Table 13.11: Multi-stage sequential diagnostic gold leach summary ...... 110 Table 13.12: Ore hardness data ...... 110 Table 13.13: The results from the chemical analyses ...... 110 Table 13.14: Test results – Master composite 1 ...... 112 Table 13.15: Natural percolation summary results ...... 112 Table 13.16: Column leach test conditions ...... 113 Table 13.17: Column leach extraction results...... 113 Table 13.18: Large column test results ...... 115 Table 13.19: Recovery of large column test ...... 115 Table 13.20: Recovery of large column test ...... 115 Table 13.21: Loaded carbon ...... 116 Table 13.22: Size by size distribution and recovery by size fraction ...... 116 Table 13.23: Reagent consumption ...... 116 Table 14.1: Drill hole database information ...... 118 Table 14.2: ATO Block model properties ...... 121 Table 14.3: Raw and composite sample statistics for Pipe 1 ...... 123 Table 14.4: Raw and composite sample statistics for Pipe 2 ...... 124 Table 14.5: Raw and composite sample statistics for Pipe 4 ...... 124 Page 9 of 209
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Table 14.6: Descriptive Statistics for Pipe 1 ...... 125 Table 14.7: Descriptive Statistics for Pipe 2 ...... 125 Table 14.8: Descriptive Statistics for Pipe 4 ...... 125 Table 14.9: Summary statistics for the capping in Pipes ...... 133 Table 14.10: Resource Metals Correlation Matrix of the ATO Deposit ...... 134 Table 14.11: Variogram Analysis Parameters ...... 135 Table 14.12: Global Mean Table for OK and ID Blocks versus composites ...... 140 Table 14.13: Mineral Resource Estimate for Oxide material in Pipe 1, 2 and 4 ...... 143 Table 14.14: Mineral Resource Estimate for Fresh material in Pipe 1, 2 and 4 ...... 144 Table 14.15: Au Equivalent Cut-Off Grade estimation parameters Mineral Resources ...... 145 Table 14.16: ATO Mineral Resource Summary for Oxide material ...... 146 Table 14.17: ATO Mineral Resource Summary for Fresh material ...... 146 Table 15.1: Pit optimization Economic Parameters ...... 148 Table 15.2: Whittle Pit Optimization Result (AuEq cut-off 0.3 g/t) ...... 150 Table 15.3: ATO Deposit Mineral Reserves for Oxide Pits ...... 154 Table 15.4: Mining dilution and Ore loss Reconciliation for ATO Deposit Mineral Reserves ...... 154 Table 15.5: ATO Deposit Mineral Reserves Statement as at August 21, 2017 ...... 155 Table 15.6: In-Pit Inferred Resources ...... 155 Table 16.1: Mine-production schedule ...... 157 Table 16.2: Mining Equipment Fleet ...... 158 Table 16.3: Contractor’s mine operating personnel Requirements ...... 159 Table 16.4: Contractor’s mine-staff requirements and salary ...... 159 Table 17.1: Design Criteria ...... 161 Table 17.2: Production plan ...... 162 Table 18.1: Water usage ...... 172 Table 18.2: Power usage ...... 173 Table 19.1: LME Historical Gold and Silver price, (USD/oz) ...... 174 Table 20.1: Details on the cost for environment ...... 175 Table 20.2: List of environmental laws relevant to placer gold and mixed metal deposit mining project...... 176 Table 20.3: List of standards ...... 177 Table 20.4 Rare and very rare plants species recorded in project area ...... 178 Table 21.1: Project investment summary ...... 188 Table 21.2: Investment per ton of ROM ...... 188 Table 21.3: Plant construction and equipment cost ...... 189 Table 21.4: Camp and infrastructures cost ...... 189 Table 21.5: Mining contractor cost per ton of rock mass ...... 190 Table 21.6: Salaried labor operating cost ...... 190 Table 21.7: Hourly labor operating cost ...... 190 Table 21.8: Processing contractor cost per ton of ROM ore ...... 191 Table 21.9: Usage of chemicals ...... 191 Table 21.10: Processing plant workers operation cost ...... 192 Table 21.11: Head office staff salary ...... 192 Table 21.12: Project operational cost and general expenses ...... 194 Table 22.1: Revenue ...... 196 Table 22.2: Non-Discounted cash flow before tax, interests, and depreciations (non-discounted) ...... 197 Table 22.3: Income statement and profit ratios ...... 198 Table 22.4: Discounted cash flow (after tax) ...... 199 Table 22.5: Sensitivity of IRR ...... 200 Table 22.6: The Cash Revenue ...... 201 Table 22.7: Profit with stream ...... 201 Table 22.8: Discounted Cash Flow with stream ...... 202 Table 22.9: Adjusted Discounted Cash Flow with stream ...... 203
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT
1. SUMMARY
1.1 Introduction
The Altan Tsagaan Ovoo Project (ATO) is located in eastern Mongolia. Independent Qualified Person Oyungerel Bayanjargal (QP as an author) from GSTATS Consulting LLC (GSTATS) and Glogex Consulting LLC (GLOGEX) were commissioned by Steppe Gold LLC. (STEPPE) to prepare a technical report to support the disclosure of the ATO’s Mineral Resource and Mineral Reserve estimate at August 21, 2017 and this has been prepared in accordance with Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101), NI 43-101 Form F1, and Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards (2014). 1.2 Property Description and Ownership
The ATO area under mining license (MV-017111) is located in the territory of Tsagaan Ovoo soum, Dornod province of Eastern Mongolia. Size of the area is 5,492.63 ha. The licensed area is located 660 km east of Ulaanbaatar city, 120 km northwest of Choibalsan, the provincial capital of Dornod province, 38 km west of Tsagaan Ovoo soum, in the Davkhariin Aryn valley and at the junction of rivers of Bayan and Duruu. It is around the foot of mountains such as Delger Ulziit, Bayan, Namkhai Hill and Yaruu. The geographic zone of ATO project is in datum WGS-84 Zone 49N of UTM coordinate system and the geographic center of the property is 48°26’N latitude and 112°46’E longitude.
As of January 31, 2017, CGM entered into definitive agreements to sell the ATO Project to STEPPE. STEPPE has paid the "rights transfer" fees to the TAX office and also paid VAT and received all required approval from Mongolian TAX office to transfer the license MV-0171111 under STEPPE. STEPPE has submitted all required paper works to MRPAM to transfer the license under its name by the end of the August 2017. The license transfer was completed on 5th of September 2017. STEPPE is now 100% owner of the mining license of ATO. 1.3 Geology and Mineralization
The geology of the ATO property consists of metamorphosed Devonian sedimentary rock overlain by a volcanic and sedimentary sequence of Permian age and remnant scraps of probable Jurassic volcanoclastic units, intruded by Jurassic plutons ranging from diorite to granite in composition and including rhyolitic phases mainly as dikes.
Mineralization on the property is probably related to the Jurassic magmatic rocks as sources of heat, metals, and fluids. Three mineralized pipes at ATO (Pipe 1, 2, and 4) have been emplaced into stratified rocks as young as presumably Early to Middle Jurassic. Pipe 4 is mainly concealed. Pipe 3 contains abundant pyrite, but no significant amounts of Au, Ag, Pb, and Zn.
The ATO deposit is an epithermal gold and polymetallic deposit of transitional sulfides in breccia pipes in a Mesozoic continental rift zone in eastern Mongolia. 1.4 Exploration and Drilling
Since acquiring the ATO Project in 2007, CGM has completed, a total of approximately 63,866 m of exploration drilling in 597 holes were completed on the project until the end of 2014. Of this, 54,425 m was core drilling in 370 holes and 9,441 m was completed in 227 RC holes. The drilling has been spread over the ATO mining license as well as south exploration area (Figure 10.1). Drilling efforts have generally focused on expanding the known mineralization, but also include confirmation drilling and exploration drilling in several
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT potential target areas on the ATO property. In addition to exploration drilling, CGM has completed geologic mapping, soil and outcrop sampling, and gravity survey at ATO. 1.5 Data Verification
CGM implemented a series of industry standard routine verifications to ensure the collection of reliable exploration data. Documented exploration procedures exist to guide most exploration tasks to ensure the consistency and reliability of exploration data. In accordance with National Instrument 43-101 guidelines, the Author QP visited the ATO deposit in August 23 and October 2, 2017. The site visit was conducted to ascertain the geological setting of the ATO Deposit gold-lead-zinc mineralization and to witness the extent of exploration work carried out on the property.
Routine verifications were completed by the Author QP to ensure the reliability of the drill hole and topo surface data, and analytical data provided by STEPPE. In the opinion of the Author QP, the electronic data of drill holes are reliable, appropriately documented and exhaustive. The analytical results are sufficiently reliable for the purpose of resource estimation. 1.6 Metallurgical Tests and Processing
Metallurgical tests for mineral processing at ATO deposit were conducted at the Central Laboratory of Xstrata Support Process in Canada, "ALS Metallurgy-Ammtec" laboratory in Australia and "Boroo Gold" LLC processing plant in Mongolia.
Mineral processing method of heap leaching was exercised based on the oxide ore metallurgical test results.
Metallurgical test samples were collected from the exploration drill core and bulk samples for both oxidized and fresh zones within mineralized pipes of Pipe 1, Pipe 2 and Pipe 3.
Metallurgical tests included a step-by-step leaching test carried out by the roll-up test (bottle-roll test) and granular ore test. Tests were conducted in 4 phases:
1. Bottle-Roll tests of the Xstrata Process Support Center in Canada (from 4 April 2010 to June 30, 2011) 2. Destruction Testing and Analysis - Canadian Xstrata Process Support (September 2011) 3. Column leach tests with low diameter (75 mm) (conducted in crushed ore) - ALS Metallurgy in AMMTEC laboratory in Perth, Australia (completed in November 2011 in April 2012) 4. Additional analysis of column tests with a large-scale diameter (280mm) - in CIL plant of Boroo Gold LLC in Mongolia - September 2012) 1.7 Mining Methods and Pit Optimization
The ATO project has been planned as an open-pit truck and shovel operation. The truck and shovel method provides reasonable cost benefits and selectivity for this type of deposit. Only open-pit mining methods are considered for mining at ATO.
Pit optimization was done to define pit limits with input for economic and slope parameters. Optimization used only Measured and Indicated Resources for oxide ore processing. All material under 0.3 g/t AuEq was considered as a waste. Whittle pit optimizations were run using the economic and slope parameters and completed using prices of $950 to $1,650 per ounce Au in increments of $65 per ounce in order to analyze the deposit’s sensitivity to gold prices. The ultimate $1,300 pit limit was determined using cash-flow analysis based on a 10% discount rate. Pit designs were created to use 2.5 meters benches for mining. Slope parameters were applied as up to 50 degree inter-ramp slope angles; up to 70 degree bench face angles. Ramps were designed to have a maximum centerline gradient of 10%. In areas where the ramps may curve Page 12 of 209
NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT along the outside of the pit, the inside gradient may be up to 11% or 12% for short distances. Design criteria accounts for 3.5 times the width of the truck for running room in areas using two-way traffic. For roads designed outside of the pit, and additional safety berm is accounted for in the road widths. The ultimate pit is an integrated design using interior pit phases and it separates 3 small mines. Based on a resource block size of 5 m by 5 m by 2.5 blocks, ore loss and dilution was assumed not more than 3% and 2%. 1.8 Recovery Methods
The recovery methods used for the oxide portion of the ATO Project will be crushing and heap leach. Flowsheet development, operating parameters and design criteria were based on results from metallurgical test work. The gold recovery process was designed on the basis of leaching 1.2 Mt of ore per year with an average gold grade of 1.13 g/t, average silver grade of 9.25 g/t at an overall gold recovery of 70%, silver recovery 40%.
The three-stage crushing plant will operate at a nominal 5860 t/d throughput, 275 days per year.
The process plant, located near and down-gradient from the heap leach facility (HLF) to minimize the pumping and pipeline requirements for pregnant and barren solutions, will operate 365 days per year. The pregnant solution will flow to the plant at a nominal rate of 200 m3/h and a design flowrate of 250 m3/h. The plant is designed to process 5 tons of carbon per day using an absorption, desorption and refining process to extract gold from the pregnant solution to produce the gold doré. 1.9 Mine Production Schedule
Proven and Probable Reserves were used to schedule mine production, and Inferred Resources inside of the pit were considered as waste. The final production schedule uses trucks and shovels as required to produce the ore to be fed into the process plant and maintain stripping requirements for each case.
Table 1.1: Mine Production schedule
Total rock Oxide ore tonnage, Waste Years Au g/t Ag g/t Au, K Oz Ag, K Oz tonnage ROM tonnage 2018-IY 544,172 302,204 241,968 1.16 8.80 11.3 86 2019 1,868,323 1,212,138 656,185 1.29 8.66 50.1 337 2020 1,909,434 1,215,459 693,975 1.29 10.45 50.3 408 2021 1,795,449 1,212,138 583,311 1.32 10.25 51.5 400 2022 1,849,953 1,212,138 637,815 1.14 11.26 44.6 439 2023 105,578 73,180 32,399 0.72 4.94 1.7 12 Total 8,072,910 5,227,257 2,845,653 1.25 10.00 209.5 1,681
1.10 Capital Costs
Investment will not be required since the open pit mining operation is planned to be carried out by a contractor. A total Investment of $ 19.59 million will be required for the whole project. Investment for ROM per ton will be $ 3.75 in ore. The investment amount includes VAT, customs duties, transportation costs and installation costs. (Table 1.2).
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Table 1.2: Project investment summary
TOTAL INVESTMENT Pre-mining overburden and Engineering design Plant construction Camp and Working capital ore stock drawings and and equipment’s Total, infrastructures, (60 days need), preparation permits expenses, cost, 1,000 USD 1,000 USD 1,000 USD expenses, 1,000 USD 1,000 USD 1,000 USD 2017-III 500 500 2017-IY 500 1,500 2,000 2018-I 500 6,007 1,000 7,507 2018-II 500 3,604 1,000 5,104 2018-III 1,201 645 1,846 2018-IY 1,317 1,201 114 2,632 Total 1,317 2,000 12,015 4,145 114 19,591
1.11 Operating Costs
The total operating cost of the project over the anticipated mine life is $ 48.81 million of which the open pit mine operating costs are $ 18.22 million, the total cost of the processing is $ 26.14 million and the general administration cost is $ 4.18 million. The total cost per ton of ore is $ 9.30 or $ 332.80 per ounce and summary shown in Table 1.3:
Table 1.3: Project operational cost and general expenses
OPEX Ore Total per OPEX Rock Mining Mining ROM Ore General General License processing operational ton per Period mass, cost, cost, ore, processing administration administration fee, Project stage cost, costs, of ounce, Tx1,000 $/t $1,000 Tx1,000 cost, $/t cost, $/t cost, $ 1,000 $1,000 $1,000 $1,000 ore, $ $ 2018 IY 544.2 2.42 - - - - 1,317 Pre -mining stage 2019 1,868.3 2.42 4,521 1,200.0 5.00 6,000 0.80 960 55 11,536 9.60 330.70 Mining stage 2020 1,909.4 2.42 4,621 1,200.0 5.00 6,000 0.80 960 55 11,636 9.70 333.60 Mining stage 2021 1,795.4 2.42 4,345 1,200.0 5.00 6,000 0.80 960 55 11,360 9.50 325.70 Mining stage 2022 1,850.0 2.42 4,477 1,200.0 5.00 6,000 0.80 960 55 11,492 9.60 329.40 Mining stage 2023-I 105.6 2.42 255 427.3 5.00 2,136 0.80 342 55 2,789 6.50 390.80 Mining stage Total 8,072.9 2.42 18,220 5,227.3 5.00 26,136 0.80 4,182 275 48,812 9.30 332.80
1.12 Revenue
The gold production is planned to begin in the first quarter of 2019 and will produce 146,669 ounces of gold and 672,518 ounces of silver during the project period. Based on market research, gold price was valued at $1306 and silver price was $21.60 per ounce. The total gross revenue of the project will be $ 206.16 million of which gold revenue is $ 191.64 million and silver revenue will be $ 14.53 million.
Table 1.4: Revenue
Gold in Silver in Gold Silver Gold Silver Gold Silver Total Gold Silver Period ROM ROM ore, produced, produced, price, price, revenue, revenue, revenue, recovery recovery ore, oz oz oz oz $/oz $/oz $ 1,000 $ 1,000 $ 1,000
2019 49,834 337,583 70% 40% 34,883.56 135,033 1306.6 21.60 45,579 2,917 48,496 2020 49,834 401,884 70% 40% 34,884 160,754 1306.6 21.6 45,579 3,472 49,051 2021 49,834 401,884 70% 40% 34,884 160,754 1306.6 21.6 45,579 3,472 49,051 2022 49,834 401,884 70% 40% 34,884 160,754 1306.6 21.6 45,579 3,472 49,051 2023-I 10,193 138,059 70% 40% 7,135 55,224 1306.6 21.6 9,322 1,193 10,515 Total 209,527 1,681,295 70% 40% 146,669 672,518 1306.6 21.6 191,638 14,526 206,164
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1.13 Mineral Reserves and Mineral Resources Summary
Table 1.5: ATO Deposit Mineral Reserve and Mineral Resource Summary (at August 21, 2017)
Notes: 1. ATO Mineral Reserves and Mineral Resources are as at August 21, 2017 using the CIM Definition Standards (2014). 2. Mineral Resources are in addition to Mineral Reserves. 3. Mineral Reserves are constrained within an optimized pit shell based on a gold price of $1,300 per ounce. 4. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability for heap leaching. 5. Contained gold estimates have not been adjusted for metallurgical recoveries. 6. Recovered gold estimates have been adjusted for metallurgical recoveries based on 70% for Au and 40% for Ag. 7. Mining dilution is 3% and Ore loss is 2%. 8. Mineral Reserves and Mineral Resources are estimated using a 0.3 g/t AuEq cut-off grade for oxide material and a 1.1 g/t AuEq.cut-off grade for fresh material. 9. A conversion factor of 31.103477 grams per ounce 453.59237 grams per pound are used in the reserve and resource estimates. 10. AuEq has been calculated using assumed metal prices ($1,306/oz for gold, $21.60/oz). Oxide ore calculation: AuEq(g/t) = Au(g/t) + (Ag(g/t) x 21.60 x 0.0321507 x 0.4) / (1,306 x 0.0321507 x 0.7) Fresh ore calculation: AuEq(g/t) = Au(g/t) + ((Ag(g/t) x 21.60 x 0.0321507 x 0.75) + (Pb% x 1,844 x 10-6 x 0.6) + (Zn% x 1,944 x 10-6 x 0.55)) / (1,306 x 0.0321507 x 0.71) 11. Inferred Mineral Resources have a great amount of uncertainty as to their existence and as to whether they can be mined economically. It cannot be assumed that all or part of the Inferred Mineral Resources will ever be upgraded to a higher category. 12. Totals may not match due to rounding.
1.14 Conclusions and Recommendations
Based on the information contained herein, the Author QP offer the following interpretations and conclusions:
Geology and Mineral Resources
ATO is an epithermal gold and polymetallic deposit of transitional sulfides in breccia pipes in a Mesozoic continental rift zone in eastern Mongolia. The procedures for drilling, sampling, sample preparation and analyses are appropriate for the type of mineralization and estimation of Mineral Resources. The classification of Mineral Resources conforms to CIM Definition Standards. Page 15 of 209
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Mineral Resources were estimated as of August 21, 2017. Combined Measured and Indicated Mineral Resources total 12.23 Mt at 1.49 g/t Au, 9.99 g/t Ag, 0.75 %Pb and 1.34% Zn containing 587 thousand oz of gold, 3,927 thousand oz of silver, 202 million lb of lead and 362 million lb of zinc. Inferred Mineral Resources total 1.05 Mt at 1.03 g/t Au, 25.18 g/t Ag, 0.52 %Pb and 1.11% Zn containing 35 thousand oz of gold, 848 thousand oz of silver, 12 million lb of lead and 26 million lb of zinc. The Mineral Resource estimate is constrained within the mineralized pipes and do not have demonstrated economic viability.
Mining and Mineral Reserves
The Mineral Reserve estimate has been prepared utilizing acceptable estimation methodologies and the classification of Proven and Probable Mineral Reserves conform to CIM Definition Standards and NI 43-101. The Mineral Reserve estimate was developed through the construction of an ultimate open pit design within the Mineral Resource model at cost estimates defined in Section 15, Table 15.1 and the reserve spot metal price assumptions of $1,306.6/oz gold and $21.6/oz silver. The Proven and Probable Mineral Reserve totals 5.23 Mt at 1.25 g/t Au and 10.00 g/t Ag containing 210 thousand oz of gold and 1,681 thousand oz of silver. The Mineral Reserve estimate takes into consideration mining and processing costs, metallurgical recoveries, sales prices, refining charges and royalties in determining economic viability. The Mineral Reserve estimate is classified as 65% Proven and 35% Probable. The methodology used for mine planning, pit limit determination, production sequence and scheduling, and estimation of equipment/manpower requirements is in line industry practice.
Mineral Processing
The metallurgical test works meet industry standards. The gold recovery process was designed on the basis of leaching 1.2 Mt of ore per year with an average gold grade of 1.13 g/t, average silver grade of 9.25 g/t at an overall gold recovery of 70%, silver recovery 40%. The three-stage crushing plant and the process plant will operate to produce the gold dore.
Economic Analysis
• Discounted cash flow calculates as discount rate at 10% per year, depreciation for 5 years, and scrap value is 40% at the end of the project and income tax is estimated 10% on under 3 billion MNT revenue (1usd=2,400MNT) up to 25% on above 3 billion MNT (or $ 1.25 million) revenue or more. • The net present value of the project is $ 67.78 million, the repayment period is 2.81 years and the IRR is 118%. • If the Revenue decreased by 20%, IRR 87% and if Investment increased 20%, IRR 99%. The IRR is 109% while operating cost increased by 20%. The ATO Project shows positive economic potential. The Author QP recommends the following:
• Assess the block model to determine if there are regions within the resource area limits, currently defined as inferred resources owing to the influence of Au assays that could potentially be upgraded to indicated and measured resources by additional drilling. Pipe 2 needs infill drilling to prove confidence and continuity. • To improve future modeling efforts, the Author QP recommends that the lithology type be less and specific for mining needs. Current lithology database has 24 different lithology codes. Some of those Page 16 of 209 NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT
lithology types were similar and combined based on the available information ended up with 10 lithology types for further analyses in this report. • Near pit exploration drilling where mineralization is open along the trend to add resource to the model. This recommendation applies to particularly Pit 4. • Continue work on refinement of bulk density estimate of mineralized material. • Conduct a significant check assay program utilizing a number of labs in addition to ALS. While standard and blank analysis results of CGM QAQC program do not raise questions relative to ALS’s results, the check assay program ensures a test of any potential bias in the existing results. This recommendation applies to production sampling as well as for exploration data. • Additional metallurgical test work is required for both Pit 1, Pit 2 and Pit 4 mineralized material. This will include bulk samples in large columns investigating crush size, agglomeration and leach rates as the mine advances.
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2 INTRODUCTION
2.1 Issuer for Whom This Report Was Prepared
This report is titled the NI43-101 Technical Report for Altan Tsagaan Ovoo Project (2017 ATO TR) and has been prepared for Steppe Gold LLC (STEPPE). 2.2 Terms of Reference and Purpose of Report
The Altan Tsagaan Ovoo (ATO) Project is located in the Eastern part of Mongolia. ATO deposit consists of series of mineralized bodies called pipes containing gold, silver, lead and zinc.
The report 2017 ATO TR was prepared for STEPPE and is based on the most current technical and cost information prepared by Glogex Consulting LLC (GLOGEX) to support project financing.
Independent Qualified Person reviewed the available studies as part of the preparation for the report.
The 2017 ATOTR presents Proven and Probable Reserves and is based on a feasibility quality level study complying with NI 43-101.
Some of the ATO deposit information and results presented herein were taken from the REPORT ON EXPLORATION WORK CONDUCTED AT ALTAN TSAGAAN OVOO PRECIOUS AND BASE METAL DEPOSIT by Centerra Gold Inc from 2012. This report is not publicly presented under NI 43-101 but has been submitted under Mongolian regulations to the Mongolian government. Annual exploration reports of CGM were used for the ATO deposit geology and previous exploration activity sections of the report. 2.3 Units of Measure and Currency
Throughout this report, measurements are in metric units and currency in United States dollars unless otherwise stated. 2.4 Sources of information and Study Participants
This report is prepared by independent QP Oyungerel Bayanjargal, acting on behalf of GSTATS Consulting LLC (GSTATS), a Chartered Professional of the Australasian Institute of Mining and Metallurgy (AusIMM), who is the author of this report. Preparation of Section 15, 16 and 22 were reviewed by Mining QP Leonid Tokar, a Professional Mining Engineer of Ontario. A detailed peer review of Section 13, 17 and parts of Section 21 relating to the process plant operating costs were performed by David Frost, FAusIMM and B. Met Eng. Section 18 and the portion of Section 21 related to Capital Cost were reviewed by Tim Fletcher, P.Eng. Dr. Martin Stapinsky, P.Geo, performed a detailed peer review of Section 20 and parts of Section 4 relating to the environmental studies, permitting, social and community aspects of the project.
The information, opinions, conclusions, and estimates presented in this report are based on the following:
• Information and technical data provided by STEPPE • Review and assessment of previous investigations • Assumptions, conditions, and qualifications as set forth in the report • Review and assessment of data, reports, and conclusions from other consulting organizations and previous property owners • A review of cost from similar operating data and performance in Mongolia These sources of information are presented throughout this report and in Section 27 of References. The QPs are unaware of any material technical data other than that presented in this Technical Report.
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2.5 Personal Site Inspections
The Qualified Person conducted two site visits on 23 August and 2 October 2017. Other visits were made in core shed and meetings in STEPPE office in Ulaanbaatar. Site visits and meetings include existing documentation and report review, fact materials investigation, geology, exploration and drilling, metallurgy, operation plans and existing development information, and previous drill hole location verification.
QPs of the report have reviewed and takes responsibility for the technical information prepared by the following individuals:
Naranbaatar Lundeg, Mineral Economist, Glogex Consulting LLC, with respect to economic matters in Sections 15 to 22. Tsogtgerel Bayarsaikhan, Mining Engineer, Glogex Consulting LLC, with respect to mining matters in Sections 15 to 22. Buyan-Ulzii Narankhuu, Mineral Processing Engineer, Glogex Consulting LLC, with respect to processing matters in Section 13 and 17. Byambatsogt Oyunchimeg, Mining Engineer, Glogex Consulting LLC, with respect to mining matters in Sections 15 to 22.
The Author QP has relied on Bataa Tumur-Ochir, CEO, Steppe Gold LLC, with respect to legal, permitting and taxation matters in Section 4.
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2.6 Abbreviations and Nomenclature
Table 2.1: List of Abbreviations
Abb. Full Description Abb. Full Description 0C degree Celsius LLDPE Linear low density polyethylene ADR Adsorption-Desorption-Recovery LME London Metal Exchange ANFO Ammonium nitrate and fuel oil m meter AR-AAS Aqua-Regia - Atomic Absorption Spectrometry m/s meter per second ATO Altan Tsagaa Ovoo Ma Millions year in age BM Block Model mbgl Meters below ground level CGI Centerra Gold Inc. mg/l milligram per liter CGM Centerra Gold Mongolia LLC MML Main Mongolian Lineament CIC Carbon-in-Column MNT Mongolian Tugrik - currency CIM The Canadian Institute of Mining, Metallurgy and N North Petroleum cm centimeter N North CV Coefficient of variation NAC North Asian Craton DCF Discounted cash flow NE Northeast D-D IP Induced Polarization Dipole-Dipole NPV Net present value DEIA Detailed Environmental Impact Assessment NW Northwest Report E East OK Ordinary Kriging EDAX Energy Dispersive X-Ray Analysis OPEX Operational Expenses EDX Energy Dispersive X-rays oz troy ounce (31.103477 g) EIA Environmental Impact Assessment pH Potential of hydrogen EPL Environmental Protection Law of Mongolia PIMA Portable Infrared Mineral Analyzer EPMA Electron Probe Micro Analysis ppb part per billion FA-AAS Fire Assay - Atomic Absorption Spectrometry ppm part per million FEL Front end loader PVC Polyvinyl chloride G&A General and Administrative QAQC Quality Assurance and Quality Control g/t gram per tone QC Quality Control GPS Global Positioning System QEMSCAN Quantitative Evaluation of Materials by Scanning Electron Microscope gr gram QP Qualified Person GRG Gravity recoverable gold RC Reverse Circulation ha hectare RL relative elevation HLP Heap leach pad ROM Run-of-mine IA Investment Agreement RQD Rock Quality Designation ICMC International Cyanide Management Code S South ICP Inductively Coupled Plasma SE Southeast ICP-MS Inductively Coupled Plasma Mass Spectrometry SEM Scanning Electron Microscope ID Inverse Distance StDEV Standard Deviation in inch SW Southwest IRR Internal rate of return t metric ton IS Intermediate Sulfidation (Sulphidation) UTM Universal Transverse Mercator k kilo (thousand) W West km kilometers WDS Wavelength dispersive spectrometry L/s Liters per second WGS World Geodetic System lb pound μ micron
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3 RELIANCE ON OTHER EXPERTS
This technical report has been prepared by the QPs at the request of STEPPE.
For the purpose of this report, the QPs has relied on information related ownership, environmental liabilities, and permits provided by STEPPE. The QPs have not researched property title or mineral rights for the ATO property and expresses no opinion as to the ownership status of the property.
The QPs have relied on STEPPE for guidance on applicable taxes, royalties, and other government levies or interests, applicable to revenue or income from ATO.
Except for the purposes legislated under provincial securities law, any use of this report by any third party is at that party’s sole of risk.
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4 PROPERTY DESCRIPTION AND LOCATION
4.1 Property Ownership and Boundaries
The ATO area under mining license (MV-017111) is located in the territory of Tsagaan Ovoo soum, Dornod province of Eastern Mongolia (Figure 1). Size of the area is 5,492.63 ha. The licensed area is located 660 km east of Ulaanbaatar city, 120 km northwest of Choibalsan, the provincial capital of Dornod province, 38 km west of Tsagaan Ovoo soum, in the Davkhariin Aryn valley and at the junction of rivers of Bayan and Duruu. It is around the foot of mountains such as Delger Ulziit, Bayan, Namkhai Hill and Yaruu (Figure 4.1). The geographic zone of ATO project is in datum WGS-84 Zone 49N of UTM coordinate system.
Figure 4.1: Location map for the ATO project
The initial number of the exploration license was 6727X when issued to Coge Gobi LLC on December 30, 2003 for the first time. The area of the license was 109,118 ha. In 2007, approximately 85,500 ha of the area were surrendered to the Cadaster Office by Order #1620 of the Head of the Cadaster Registration Center, leaving Page 22 of 209
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23,597.51 ha to the license. This reduced license was then transferred in full to Centerra Gold Mongolia LLC (CGM) by Order #513 of Head of the Cadaster Office of the Mineral Resources Authority dated May 4, 2010. In Mongolian regulations, exploration license owner should submit exploration result reports every year. CGM conducted intense exploration programs since 2010 and reduced license area based on the exploration results.
CGM prepared a Feasibility Study report based on the Mongolian guidance in 2012 to transfer the exploration license to a mining license. MV-017111 mining license was issued to CGM with area of 11,614.22 ha on August 31, 2012 and extended for 30-year term expiring on August 31, 2042. The license was surrounded by 19 location points that time. Now, MV-017111 license boundary was simplified to 8 points with area of 5,492.63 ha and shown in Table 4.1 and Figure 4.1.
Table 4.1: ATO Project license boundary coordinates
Name and number of Area (ha) Coordinates licensed area x y 112° 41' 32.1" 48° 24' 35" 112° 41' 32.1" 48° 27' 27" Urd Tsagaan Ovoo 5,492.63 112° 45' 43" 48° 27' 27" 112° 45' 43" 48° 28' 14" MV-017111 112° 49' 20" 48° 28' 14" 112° 49' 20" 48° 26' 44" 112° 48' 48" 48° 26' 44" 112° 48' 48" 48° 24' 35"
As of January 31, 2017, CGM entered into definitive agreements to sell its 100% interest in the ATO Project to STEPPE for $ 20 million. STEPPE has paid $ 0.2 million for the "rights transfer" fees to the TAX office including VAT and received all required approval from Mongolian TAX office to transfer the license MV- 0171111 under STEPPE. STEPPE has submitted all required paper works to MRPAM to transfer the license under its name by the end of the August 2017 and the $ 9.8 million was paid to CGM for license transfer. The mining license was transferred to STEPPE on 5th of September 2017 by Decision #544 of Head of Cadaster Office of the Mineral Resources and Petroleum Authority. STEPPE is now 100% owner of the mining license of the ATO. Remaining $ 10 million will be paid within first two year of mining period from STEPPE to CGM. 4.2 Investment Agreement (IA)
At the time of signing IA with Mongolian government, a company must submit several reports such as updated Mineral Resource and Mineral Reserve report, Feasibility Report, and Environmental Impact Assessment Report under Mongolian guidance. 4.3 Royalties
The royalties applied to the ATO mine operation is a Mongolian Government State royalty according to the Article 47 of the Minerals Law of Mongolia (The Minerals Low). The Clause 47.3.2 of the Minerals Law demonstrates minimum of 5% of the final production selling cost be paid if metals other than gold sold to the Mongol Bank or a bank credited by Mongol bank and additional royalty percent be added depending on the market metals price at the time of the selling sliding scale demonstrated in the Clause 47.3.5 of the Minerals Law. For the gold, Clause 47.3.3 demonstrates that if gold sold to the Mongol Bank, royalty will be 2.5% and additional percent indicated in Clause 47.3.5 will be 0%.
The sliding scale in Clause 47.3.5 of the Minerals Law for gold, silver, lead and zinc is shown in Table 4.2 and 4.3.
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Table 4.2: Mongolian Government State sliding scale royalty for Au and Ag
Gold price ($/oz) Royalty Silver ($/oz) Royalty 0-900 0% 0-25 0% 900-1000 1.0% 25-30 1.0% 1000-1100 2.0% 30-35 2.0% 1100-1200 3.0% 35-40 3.0% 1200-1300 4.0% 40-45 4.0% 1300 and above 5.0% 45 and above 5.0%
Table 4.3: Mongolian Government State sliding scale royalty for Pb and Zn
Royalty Lead price ($/t) Ore Concentration Production 0-1500 0% 0% 0% 1500-1800 1.0% 0.8% 0.4% 1800-2100 2.0% 1.6% 0.8% 2100-2400 3.0% 2.4% 1.2% 2400-2700 4.0% 3.2% 1.6% 2700 and above 5.0% 4.0% 2.0% Royalty Zinc ($/t) Ore Concentration Production 1-1500 0% 0% 0% 1500-2000 1.0% 0.8% 0.4% 2000-2500 2.0% 1.6% 0.8% 2500-3000 3.0% 2.4% 1.2% 3000-3500 4.0% 3.2% 1.6% 3500 and above 5.0% 4.0% 2.0%
4.4 Required Permits
Approval for commencement of mining operation
The license holder is entitled to obtain permission to commence mining operations in accordance with Clause 35.4 of Mineral Law after completion of the Updated Resource report, Updated Feasibility Study, Updated Detailed Environmental Impact Assessment and the construction of the mine. STEPPE needs to complete those updated reports prior to potentially initiating production at ATO in October 2018. The Author QP suggests that those update reports need to be completed as early as possible.
Environmental obligations of the mining license holder
Under the Environmental Protection Law of Mongolia (the “EPL”), business entities and organizations have the following duties, including, but not limited to, with respect to environmental protection:
(a) to comply with the EPL and the decisions of the government, local self-governing organizations, local governors and Mongolian state inspectors; (b) to comply with environmental standards, limits, legislation and procedures and to supervise their implementation within their organization; (c) to keep records on toxic substances, adverse impacts, and waste discharged into the environment; and (d) to report on measures taken to reduce or eliminate toxic chemicals, adverse impacts, and waste.
The Minerals Law and the Law on Environmental Impact Assessment (EIA) state that the mining license holder shall prepare and submit an environmental impact assessment report to relevant government authority for its approval.
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According to the procedure for the Feasibility Study for the Use of Mineral Deposit by the Minister of Mineral Resources and Energy dated April 17, 2012, the ATO Feasibility report should be updated. Then a detailed environmental assessment report is required to be updated after the approval of a feasibility report by Professional Committee of Mineral Resource and Reserve. The ATO Feasibility Report update is expected to be completed by April 2018.
Mining license holder must annually develop and implement an environmental management plan (EMP) which includes also environment protection plan to be submitted to the relevant authorities. Accordingly, the environmental management plan shall be approved (Clause 39.1.7 of Minerals Law).
Other relevant permits
(a) Land possession right
Mining license holder can request to use or possess land for the mine claim, camp and other permanent facilities for its mining operations.
Only Mongolian citizens and legal entities are entitled to hold a land possession right (Clause 32.1 of Land Law).
Land use or possession rights are evidenced by a Land Use or Possession Certificate issued by a local authority of the soum (district) in which the relevant land is located. The terms and conditions of land use or possession rights are governed by a land possession agreement entered into with the land division of a local authority of the relevant soum (district).
The land use agreement will be signed by the end of October, 2017.
(b) Water Use Permit:
Mining license holder shall obtain Water Use Inference from State central administrative organization or local authority depending on proposed amount of water usage. License holder shall submit a request for Water Use Permit to an authority granted the Water Use Inference. Upon the issuance of Water Use Permit, Water Use Certificate is granted and Water Use Agreement is executed. Transferring water use permit is prohibited.
As informed by the STEPPE official, Water Use Permit is not obtained yet. Water permit shall be obtained from the local governor and the water authority for industrial quantities.
(c) Energy and Powerline Permits
Power line permits will not be obtained during the mining of the oxide ore. At this stage, the power will be supplied by a diesel generator.
(d) Construction Permit
The ATO project requires permits to construct the following building units. STEPPE plans to obtain the necessary permits and complete construction within the following periods:
Camp by October 2017 Water supply system by April 2018 Heap leach facility by April 2018 Tailings pond by April 2018 Processing plant and laboratory by April 2018 Diesel generator foundation by April 2018 Warehouse and fuel storage by April 2018
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Maintenance, workshop and administration office by April 2018
(e) Explosive Permit
Blasting will be performed by a blasting contractor that will have the required permit. 4.5 Socio-Economic Impacts
Centerra Gold Inc. (CGI) is a parent company of CGM assets in Mongolia and owns a 100% interest in the Boroo Mine, a 100% interest in the Gatsuurt property, and was a 100% interest in the ATO property. Boroo is an open-pit operation and Mongolia’s largest hard-rock gold mine. It began commercial production in 2004 and produced approximately 1.8 million ounces of gold from March 2004 through to December 2016. Boroo is operated through CGI’s Mongolian subsidiary, Boroo Gold Company. The operations at the Gatsuurt property are pending due to suspended negotiations with the Mongolian Government. The Project will begin ore production on receipt of the final approvals and permits (Centerra Gold Inc. 2011a). The Gatsuurt property is being currently advanced by CGM, an exploration group completely owned by CGI. CGM owned 100% interest in the ATO Project which was later acquired by STEPPE in 2017.
At this stage of the ATO Project, the socio-economic management system of STEPPE is currently being developed. Presumably, it will evolve along the same lines, and drawing on the same principles, as that of Boroo Gold Company. A major pre-condition for this is that STEPPE sets out corporate environmental, social, economic and ethical policies that direct and regulate all its performance. 4.6 Community Policy, Planning, Consultation and Engagement
This Social Baseline Assessment has provided useful preliminary information for the planning of Project developments by CGM and it will provide STEPPE with an effective pathway for conducting additional public consultations with local communities. The Author QP is not aware of any environmental liability or significant issues to affect development of the property. Careful Project planning, including the implementation of an international-style social and environmental impact assessment (SEIA), a Mongolian EIA, and a land use management and reclamation plan, will be important components for the success of the project.
There are several key observations drawn from this Social Baseline Assessment, based on survey work in the field, research and analysis, and also from the expressions of concern and interest by the local people. These observations include:
• There are archeological sites on the ATO LA and sacred places located in the general area: • sacred places must be fully avoided and protected from any sort of Project activities; • archeological sites should also be avoided. If absolutely necessary to use the area for mining and/or infrastructure, archeological excavations should be planned and conducted prior to the mining of the site. • New jobs will be created that will provide opportunities for local people, including possibly youths. The local people believe that: • use of local labor and the conditions for initial training should be investigated by STEPPE; • a hiring policy that optimizes the opportunity for hiring a local labor force should be adopted by STEPPE, and proper training programs should be developed. • Regular communications with local authorities and local people will be vital to the success of the Project. To address this important aspect of the Project, it is recommended that STEPPE should: • establish a local community liaison office; • conduct regular communications with all stakeholders with particular attention paid to vulnerable groups (nomadic communities, retired people, others); • develop mutually acceptable communication rules with the local people. • Participation/sponsorship in local infrastructure development and community development projects by STEPPE, as well as a safe driving program, are important issues for the local people.
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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY
5.1 Topography, Elevation, Vegetation and Wild life
Geographically, the license area is located in the low mountain zone at the north-east end of the Khentii Mountain Range and at the south-west part of the Dornod high steppe. The topography of the project area generally consists of small rounded, separate mountain complexes with small hillocks in a steppe. Average elevation is 980-1050 m above sea level, with the lowest point being Deliin Well (979.3 m) and the highest point being Mount Temdegt (1144.7 m). Relative elevation is 60-120 m. Mostly, brown and black brown gravel, sandy loam, and gravel-mild clay of steppe zone are predominant.
Quaternary loose sediments are widespread in the license area where there are no trees. Mounds and hills in the area are bald and rounded. Vegetation and grass cover the entire area and includes pasture plants such as khazaar grass, wormwood, stipa, brome-grass, and couch grass.
Hoofed animals of steppe in the region include white gazelle. Carnivores are wolf, fox and corsac. Rodents include marmot, gopher, shrew-mouse, and stoat. Birds include lark, red nose, crane, bustard, scoter, and brown nose. Also, crawlers, locust, grasshoppers, mosquitoes and midges are abundant. 5.2 Property Access
The ATO property is situated in the western side of Mongolia, 660 km east of the city of Ulaanbaatar, 120 km northwest of Choibalsan city and 38 km west of Tsagaan Ovoo soum. Access to the property from the Mongolian capital, Ulaanbaatar is possible by highway to Choibalsan, the centre of the Dornod province and improved unpaved road continue to Tsagaan Ovoo soum. The property is connected to other settlements by dirt roads. It is possible to fly to Choibalsan from Ulaanbaatar via domestic airlines. 5.3 Regional Population and Infrastructure
The nearest settlement to the property is the central village of the Tsagaan Ovoo soum, which is settled at side of the Khuuvur Lake, with moderately developed infrastructure. The Tsagaan Ovoo soum consists of six subsections and has total population of 3800. Nationality consists of 80% Buryats and the rest of Khalkh people. The community is mainly engages in domestic animal husbandry. Some of them run plantation and grow vegetable for their own household uses. The central village accommodates administrative offices, a cultural center, secondary schools, a hospital, a kindergarten, a communications center, cellphone stations, a gas station, and high-voltage sub-stations. 5.4 Climate and Operating Season
ATO mine will operate all year around. Climate of the property region is characterized by extreme cold and hot weather like Boroo mine site. Daily, monthly, and yearly fluctuations of temperature are common. Winter is harsh and cold and extends from November to March. Coldest temperature reaches to -460C near the river valleys and flatlands in January. Stable snow cover persists from November to March. Freezing of soil starts from mid-September and continues till late May, with the freezing depth reaching 2.5 m. Spring begins in late March and continues till early June, and it is characterized by fluctuant atmospheric temperature and air dryness and strong wind. Summer is shorter than other seasons, dry and chilly. The hottest temperature is up to +400C in summer. 60-80% of the annual precipitation falls as rain during July and August. Number of days with precipitation is 59 days in average in a year. Average wind speed is 4-8
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT m/s, with dominant directions from north-west, north and north-east and maximum speed reaches 2022 m/s. (Source: National Atlas of Mongolia, 2009) 5.5 Water
Water network of the area belongs to the Pacific Ocean basin. Small local rivers with short-lived stream fed from eastern branch mountains of the Khentii Range flow into small lakes. Sizes of these rivers vary depending on their main source of water collection which is precipitation. Drinking water could only be taken from wells due to low density of water network. Lakes in the region such as Duut, Tsagaan, Ovoot, Eregtseg, Ukhaagiin Tsagaan, Davkhariin Tsagaan, and Khaichiin and many other small salt lakes are also fed by rainfall. In recent years, small rivers and streams have been dried up due to a global warming and decrease of precipitation. In the summer time, seasonal springs found from the melt of small patchy permafrost in small intermountain valleys and from seasonal thawing of frozen ground. 5.6 Land Use
The land surrounding the property is predominantly used for nomadic herding of goats, cows, horses and sheep by small family units. Use is based on informal traditional Mongolian principles of shared grazing rights with limited land tenure for semi-permanent winter shelters and other improvements. 5.7 Risk Assessments
The Law on Environmental Impact Assessment (2012) and the guidelines require the inclusion of a risk assessment in project documentation. This means identification and prediction of the possible emergencies and accidents that could occur during the production process or natural disasters, and elimination and mitigation of their consequences.
There are mining license for mining operations and related permits, the availability of power source, water, potential areas for waste material and heap leach pad area, mining personnel and local working power in the project area. 5.8 Closure and Reclamation
As part of overall project planning, a preliminary reclamation and closure plan was prepared. Certain features of the mine, such as the open pit, waste dumps, and tailings impoundment, will create permanent changes to the current landscape that cannot be completely remedied through reclamation. The closure plan will; however, ensure that these disturbed areas are seismically and chemically stable as to limit the ecological impacts to the surrounding water, air, and land.
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6 HISTORY
6.1 History and Land Holdings
The area of immediate interest at ATO is centered on three, at most 200 m high, gently rounded hills (herein termed Hills 1, 2 and 3 from SE to NW) that are surrounded by widespread unconsolidated Quaternary deposits (Figure 7.3). Presently, the known most intensely mineralized rocks underlie Hills 1 and 2, respectively underlain by Pipes 1 and 2, as well as the completely covered Pipe 4. Low gently rounded slopes or somewhat steeper inclines with moderate slopes characterize the surrounding ranges.
Regional geologic field parties working for COGEGOBI were the first to recognize mineralized rocks cropping out at ATO. In 1997 COGEGOBI (a wholly owned subsidiary of the French multinational company AREVA) began their exploration efforts in eastern Mongolia. Finally, after a six year long reconnaissance effort, COGEGOBI in 2003 settled on a selected region, and then obtained eight exploration licenses in eastern and southeastern Mongolia. COGEGOBI then embarked upon a four-year-long concerted effort in the search for viable Au and uranium deposits in all eight of their licensed areas. Two of their licenses (3,425.5 km2 in all) were in the general area of ATO. As part of that effort, COGEGOBI geologists eventually collected 52 grab samples from outcrops and subcrop at ATO, and identified anomalous gold concentrations in vein quartz- rich rock there (0.06 to 27.8 g/t Au). They initially were led to the occurrence by a stream sediment sampling program. Almost all Au–anomalous samples collected by COGEGOBI are from Hills 1 and 2. COGEGOBI then proceeded to describe the occurrence at ATO as “intense hydrothermal alteration associated with volcano- plutonic structures” (Hocquet, 2005).
However, at the end of a four-year-long, exploration cycle by COGEGOBI in their eight licensed areas, which was primarily focused on gold deposits but also included uranium, uranium prices in mid-2007 skyrocketed from $30US per lb. U3O8 to about $140US per lb. U3O8. AREVA subsequently acquired East Asia Minerals Energy Company in September 2007, became AREVA Mongol in 2008, and now (2017) has a 100 percent interest in COGEGOBI. Consequently, AREVA Mongol’s interest shifted from precious metals to the energy sector, and one of their two exploration licenses surrounding ATO was dropped.
In addition, COGEGOBI geologists probably were discouraged at the economic potential of ATO by their belief that all high gold-bearing rocks resided in a small volume (i.e., low tonnage) of near-surface mineralized quartz veins above a shallow-dipping contact with underlying low-to-nil gold-bearing rocks (Hocquet, 2005). Subsequent exploration efforts at ATO by CGM proved this concept incorrect.
Initial Field Examination by CGM and Discovery
CGM geologists were invited by AREVA to visit ATO in the fall of 2009. They were accompanied during their field examination by AREVA Mongol geologists. Subsequently, initial agreement to explore the one remaining licensed area (Tsagaan Ovoo) encompassing ATO was obtained from AREVA Mongol by CGM in May 2010, and all rights to the property were obtained by CGM late in 2010. AREVA retains a 1.75 percent NSR in the Tsagaan Ovoo license.
Key outcrops in the ATO property are at Hill 1 where gold-bearing silica sinter is now considered to represent a paleo hot spring environment, previously, to the best the knowledge, unknown in Mongolia. It forms a silica cap to underlying base and precious metal mineralized rock, basically at a Mesozoic paleo surface. Reticulated mats of silicified reeds, feathery in longitudinal section and oval in cross section, are well preserved in this sinter at ATO (see below). The sinter is highly phosphatic, > 1,500 ppm phosphorous, all of which calls to mind features forming today at the surface at the Yellowstone geyser field.
But most importantly, siliceous sinter at ATO includes free gold. Silica cap includes mostly quartz (sparse chalcedonic fabrics are preserved; bladed silica locally replaces calcite) and variable concentrations of iron Page 29 of 209
NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT oxide minerals. Less abundant are kaolinite, illite, numerous primary and a number of secondary Pb minerals. Secondary Pb minerals include Pb–Al phosphate plumbogummite; pyromorphite (also a Pb phosphate); and Pb–Mn oxide minerals. Sparse galena in siliceous sinter was encapsulated in quartz, as is rare sphalerite, arsenopyrite, and argentite. Barite was present in narrow micro veins, as well as large tabular crystals in open cavities.
Even during CGM’s first brief visit to the property, it became clear that exposures of the mineralized system at ATO include highly disrupted, near paleosurface epithermal, silica-dominated rocks that had the potential to host a significant tonnage of precious- and base-metal mineralized rock. Four grab samples collected by CGM (1.3 to 3.3 g/t Au), during its initial visit to the property, confirmed presence of the anomalous Au originally reported at the ATO site by COGEGOBI.
The recent exploration history at ATO, thus, comprises three stages: (1) initial work by COGEGOBI/AREVA, (2) a due diligence stage by CGM that also included soil geochemistry and limited IP, and (3) post May 2010 comprehensive exploration by CGM. Beginning in May 2010 and through December 2014, CGM designed significant amount of exploration work in the area, incorporating geologic mapping (including ASTER imaging), widespread grab sampling, additional grid soil sampling, stream-sediment sampling, geophysical surveys (air mag, ground mag, IP, gravity), trenching, and extensive core drilling. In addition, a wide-ranging district-wide grab sampling program was conducted in association with regional geologic mapping.
As a result of these efforts, it became readily apparent that the bulk of the presently known precious and base metal mineralization at ATO is in three carrot-shaped vertically downward plunging, presumably Jurassic breccia pipes. Two of the pipes (Pipes 1 and 2) respectively underlie Hills 1 and 2, and Pipe 4 is completely concealed to the SE of Pipe 2 (Figure 6.1). 6.2 Previous Mineral Reserves and Mineral Resources
Centerra Gold Inc. reported Mineral Resource summary at ATO in its 2016 Annual Information Form on SEDAR in May 31, 2017 and shown in Figure 6.1. The information is relevant and reliable and has been superseded by the Mineral Resource estimate in Section 14 of this report.
Figure 6.1: ATO Mineral Resources in 2016 AIF by Centerra Gold Inc. SEDAR
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Recent exploration report on ATO project was prepared in 2012 for Mongolian jurisdiction. That report was not prepared under compliance of NI 43-101 and authors of such report appear not to have been qualified. Although that report was not compliance under NI 43-101 most of the information, studies prepared similar guidelines or standards that were acceptable to the NI 43-101.
There has been no production from the property.
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7 GEOLOGIAL SETTING AND MINERALIZATION
7.1 Regional Geological Setting
Relative to its continental-scale geologic framework, ATO is situated within the Devonian through Late Jurassic Mongol-Okhotsk tectonic collage (Figure 7.1) that has been emplaced along a transform-continental margin of the North Asian Craton (NAC) as shown by Parfenov and others (2010).
In addition, the Transbaikalian-Daxinganling transpressional magmatic arc that is present south of ATO along an ENE, 2,000 km long trend was thought to range in age from 175 to 96 Ma (Middle Jurassic to Early Cretaceous) (Parfenov and others, 2010).
Mongol-Okhotsk Belt
Onon base metal province
Figure 7.1: Location of the Mongol-Okhotsk belt and Onon precious-base-metal province residing between the Siberian and North China Plates (CGM 2012 ER)
Regional metallogenic setting of ATO is important from an exploration perspective. Mineral deposits range widely in age throughout Eastern Mongolia and neighboring regions of Russia and China. For example, the China Altay hosts 380–360 Ma siliciclastic VMS deposits with bimodal geochemistry in a major magmatic arc (Goldfarb and others, 2003). Further, as summarized by Xiao and others (2009), end-Permian to mid-Triassic docking of the Tarim and North China cratons against the Siberian craton resulted in (1) closure of the Paleoasian Ocean, and led to (2) formation of a number of world-class metal deposits, some of which are Triassic in age.
A number of Late Jurassic-early Cretaceous (175–96 Ma) broad, gold-bearing mineral belts also have been recognized in eastern Mongolia and in the surrounding region (Rodionov and others, 2004). Their Middle Jurassic-Early Cretaceous (175–96 Ma) time slice yields 31 gold-bearing mineral belts among 56 belts in all (55 percent)—the most mineral belts outlined for the various time slices established in the above-cited report.
Most gold-bearing belts during the Middle Jurassic-Early Cretaceous have moved decidedly “inboard” towards the Siberian craton relative to older belts, and they are present in China and Mongolia, as well as eastern Siberia. Areal distribution of the gold-bearing belts generally follows the tectonic grain or trend of the various geologic terranes and their overlap and “stitch” assemblages throughout the region.
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ATO is located north of the Main Mongolian Lineament (MML – red dashed line in Figure 7.1) and midway along the NNE trending, 600km long Onon base and precious-metal province that crosses eastern Mongolia (Figure 7.1).
The overwhelming bulk of the Pb–Zn occurrences and deposits in eastern Mongolia are located north of the MML and east of the Onon trend. The Novo and Lugiin polymetallic deposits in the Russian Federation were used to anchor the northern terminus of the Onon province as depicted in the Russian Federation (Figure 7.1). The two major mineralized trends or metallotects in this part of Mongolia (Onon and Yeroogol, red dashed line) parallel Lake Baikal, and must represent deep-seated splays possibly dating from zones of crustal weakness first developed at the time of Devonian-age accretion and dislocations along the MML. However, rifting in Lake Baikal is much younger as it began about 30 m.y. ago; i.e., during the Middle Oligocene.
Therefore, the two mineralized trends (Onon and Yeroogol) must mark zones of rifting in the earth’s crust, somewhat deeper seated than that at Lake Baikal, and whose regional least principal stress direction must have been oriented NW–SE (present day coordinates). However, this orientation of the regional least principal stress must represent a clockwise rotation from its essentially EW orientation that prevailed during final stages of rifting in the Cretaceous following mineralization and associated magmatism at ATO (see below).
A number of Hg–Sb occurrences are aligned closely along the trace of the Onon trend, as are numerous clusters of gold occurrences that are associated somewhat more broadly in the immediate region.
Further, as defined herein, the Onon base and precious-metal province coincides with relatively thick crust, roughly greater than 125 km in thickness, which partly explains abundance of base metals (especially Pb) throughout the province. Rocks south of the MML are considered to be largely Silurian to Carboniferous island-arc volcanostratigraphic packages of rock.
Though ATO presently represents the only well-explored gold deposit in this part of Mongolia, a large number of minor gold occurrences have been recognized throughout the region. Most of these gold occurrences are located outside the Onon province, as are the overwhelming number of recognized porphyry systems. The Onon province also includes a number of primarily Ag occurrences, as well as a few porphyry systems, including the Avdartolgoi porphyry, about 200 km NE of ATO, which only is mentioned briefly in passing by Dejidmaa and others (1999) without any further details.
A number of major base and precious-metal deposits also are present in the region, a metal association similar to ATO. These include the Novo and Lugiin polymetallic deposits at the northern distal end of the Onon province (Figure 7.1). 7.2 District Geology and Magmatism
To use the term “district” in its broadest sense (“sensu lato”), because no unifying genetic model or linked group of models has been established confidently for all mineralized occurrences close to ATO other than a presumed association with Jurassic magmatism. Further, the age of gold-mineralized rock has not been determined radiometrically either at ATO, or at mineralized rock in most all its surrounding occurrences.
The oldest layered rocks in the ATO district are Devonian. Relatively small isolated areas of outcrop of Early Devonian trachyrhyolite, trachyrhyolite porphyry, ignimbrite (welded tuff), and minor limestone are present in the northern part of the ATO district (Figure 7.2). The Devonian rocks are in tectonic contact with Early and Late (?) Permian strata along pre-Lower Cretaceous NW–striking, high angle faults near the NW corner of the area, and Lower Cretaceous rocks overlie unconformably the Devonian rocks. Devonian rocks are intruded by Early Permian leucogranite near the NE corner of the district (Figure 7.2).
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Though the ATO district includes limited exposures of Devonian and Triassic rocks, the most widespread rocks in the district are Early Permian volcaniclastics including tuff breccias, as well as high K andesite, and rhyolite exposed in the cores of broad uplifts. Zircons from rhyolite have been dated at 285.9 Ma (Early Permian) by a major ongoing collaborative program by CGM with Jim Mortensen at the University of British Columbia to date magmatism and mineralization in the district. The Early Permian volcaniclastics in the ATO district are further intruded by early Permian leucogranite, plagiogranite, and diorite; zircons from plagiogranite have been dated at 279.5 Ma.
Nonetheless, Permian strata largely form the cores of broad horsts in three areas: (1) a relatively small area of outcrop (2 km in long dimension) near the NE corner of the district; (2) an approximately 25 km long, NW elongated expanse that extends from the east central part of the area to the north edge of the district; and (3) a 16 km long, NS elongated belt that extends from about 5 km N of ATO to the south edge of the district. Rocks shown as Early Permian strata are mostly volcanic affiliated (rhyolite, dacite, ignimbrite, andesite- dacite porphyry, tuff, and tuffaceous sandstone), whereas strata assigned provisionally to the Late Permian are mostly sandstone, conglomerate, siltstone, and tuff.
Early Permian diorite crops out near the NE corner of the district (Figure 7.2). The age of this unit is inferred from the age of rocks that intrude it. The diorite in this area is intruded by Late Early Permian leucogranite, plagiogranite, and granodiorite, as well as Early Jurassic granite and granodiorite at the Bayan Munkh prospect. Late Early Permian leucogranite, plagiogranite, and granodiorite crop out mainly in four areas: (1) near the NE corner of the area; (2) near the west-central edge of the area, (3) NW of ATO, and (4) near the south edge of the mapped area. Early Permian leucogranite intrudes Devonian rocks as well as rocks assigned to both Permian units (P1 and P2).
At ATO, presumably Middle-Late Jurassic (?) gravel and coarse pebbly sandstone, as described above, fill a broad shallow depression on the flanks of a Permian-cored uplift (Figure 7.2). Mesozoic intrusive rocks also crop out in various locales in the district.
Some are mineralized as at Bayan Munkh. Cretaceous sedimentary rock unconformably rests on all of the older units in the district (Figure 7.2). Cretaceous siltstone, sandstone, conglomerate, and basalt fill a narrow NS Cretaceous graben NW of ATO. Preliminary PIMA examination of four samples of laminated siltstone in a drill hole into the graben indicate presence of chlorite (D. John, written commun., 2012), as opposed to presence of mixed layer clays now known to form the host mineral for lithium in similar basins elsewhere. A number of prominent NS striking faults pass just to the E of ATO, and NE-striking, high-angle faults also are present at ATO.
The Cretaceous graben reaches its maximum width of about 6 km approximately 22 km N of ATO. Coarse- grained Paleozoic granite (Early Permian?) also is well exposed and widespread west of the graben that probably opened in response to crustal EW (present coordinates) regional extension beginning in the Late Cretaceous.
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Figure 7.2: District geological map of ATO project
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7.2.1 Mineralization in the District Surrounding ATO Though the ATO Project is the most important discovery to date and is present in the south-central part of the district, a number of other mineralized occurrences also are present nearby including from N to S the Bayan Munkh, the High Land, Duut Nuur, Bayan Gol, Mungu, Apricot and Davkhar Tolgoi prospects. CGM discovered in the exploration area a hard-rock gold-lead-zinc deposit and similar occurrences and mineralization points as well as two gold occurrences, four gold mineralized points, three lead mineralized points, three lead-zinc mineralized points, and several secondary dispersion halos of gold, lead, zinc and silver. 7.3 Geology of the ATO Deposit
7.3.1 Geology Three mineralized pipes at ATO (Pipe 1, 2, and 4) have been emplaced into stratified rocks as young as presumably Early to Middle Jurassic (Figure 7.3 and 7.4). Pipe 4 is mainly concealed. Pipe 3 contains abundant pyrite, but no significant amounts of Au, Ag, Pb, and Zn. There is a strong Au anomaly in soil at ATO, as well as other accompanying metals, particularly Pb. In the aeromagnetic field, only post-mineral young dikes have a prominent positive response; much weaker responses outline some ring-shaped features. ATO resides near the center of the latter. Sinter and silicified rocks are reflected as shallow resistivity anomalies, but broad clay and chlorite-altered rocks are characterized by low resistivity. The pipes coincide with chargeability anomalies that overall really are quite weak.
Figure 7.3: Mineralized pipes at ATO deposit (View to WNW)
Figure 7.4: Plan map of ATO mineralized pipes.
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Surface projection of the morphology of Pipes 1, 2, and 4 is shown in Figure 7.4. An upper zone of Au-Pb-Zn- Ag mineralized rock at Pipe 1 is approximately oval in shape and is about 320 m wide. Pipe 2 is elongate to the NE and is approximately 320 m by 160 m in maximum dimension. Pipe 4, which is completely concealed, also is elongate to the NE and is about 400 m by 200 m. As will be shown in the cross section below along section line A-A’ through the pipes, the pipes taper slightly with depth and have a carrot-shaped 3D configuration, narrowing gradually to depths of approximately 200 m.
The deepest hole into Pipe 1, inclined 60 degrees, is about 700 m. Silica cap rock has a variable thickness in Pipe 1, generally tapering from a maximum thickness of about 40 m under the topographic high point of the pipe to less than 1 m near its margins. However, bottom surface of the cap rock is highly irregular, showing sharp undulations with underlying quartz-veined Middle-Late Jurassic (?) gravel and coarse pebbly sandstone, some blocks of which are totally engulfed by massive silica.
Figure 7.5: Longitudinal cross section through Pipe 1 and Pipe 2. Au and Ag grades in right and left sides respectively.
The pipes also are cut by a number of minor faults, both steeply dipping and shallow dipping. Some narrow flat-lying post-mineral diorite dikes also have been emplaced along faults that offset margins of the pipes.
Pebbly conglomerate and pebbly sandstone were being shed from both nearby mostly Early Permian highlands elevated during emplacement of Early Jurassic magmatic rocks, as well as apparent high walls of an enclosing oval collapse feature. Continued deposition of Jurassic strata then covered the pipes after cessation of mineralization.
Despite the three mineralized pipes being so geographically close to one another, there are distinct differences in their metal geochemistry. Pipe 2 is notably base-metal enriched; Pipe 1 contains less base metals, and Pipe 4 contains further decreases in base metals, particularly near its margins where extremely high Ag contents (locally 100s of ppm Ag across narrow intercepts) are present in association with base metal concentrations of a few hundred ppm in all. 7.3.2 Controlling Factors Important geologic controlling factors for the mineralized pipes at ATO include:
(1) Presence in a major base metal (Pb-Zn) province (thick crust) known to include large tonnage Au-Pb-Zn deposits, combined with
(2) Near paleo surface, epithermal (hot spring) emplacement of the upper parts of mineralized pipes associated with Jurassic magmatism into a near surface
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(3) Shallow depression where Middle-Late Jurassic (?) pebbly conglomerate and pebbly sandstone were being deposited prior to mineralization and continuing after mineralization.
Origin of brecciation in the mineralized pipes remains unclear, however, and shows features of both magmatic and hydromagmatic breccias (see Sillitoe, 1985). Some fine-grained, matrix-supported breccias with abundant rock flour encountered at depth at ATO are typical of a magmatic breccia-style diatremes (magmatic hydrothermal systems that extend to surface). 7.3.3 Silicate Alteration in Mineralized Pipes ATO is characterized by an absence of adularia that is relatively abundant elsewhere in epithermal Au-Ag deposits. As discussed below, at ATO this primarily is a reflection of high Mg/2H+ ratios and a correspondingly low K/H+ ratio in mineralizing fluids associated with an underlying largely dioritic magmatic complex. At ATO, siliceous sinter forms a cap rock at the top of Pipe 1 and includes barite (in narrow micro veins as well as tabular crystals in open cavity fillings), clinochlore (Mg chlorite), and less abundant Mg-Fe chlorite (Figure 7.6 and 7.7). With increasing depth, silica is increasingly present in the pipes as relatively late paragenetic stage vein quartz filling central parts of sulfide mineral-rich veins. Though clays (mostly kaolinite) occur with sinter at the top of the mineralized column at Pipe 1, clinochlore (Mg chl) is the dominant alteration hyrosilicate mineral at depth throughout mineralized breccia. At depths near 200 m, however, clinochlore begins to give way to phlogopitic white mica. White mica also increases near margins of Pipe 4. Gypsum (after anhydrite) is concentrated near margins of the pipes.
Figure 7.6: Schematic cross section through Pipe 1 at ATO (A) showing relative abundance of alteration minerals with depth (B)
Pipe 2 at the surface, in place of a massive silica cap, instead is marked by variable concentrations of quartz veins and veinlets in networks that cut Middle-Late Jurassic (?) pebbly conglomerate and pebbly sandstone. The veined pebbly conglomerate and pebbly sandstone at this pipe extends outward to where it eventually is covered by unconsolidated Quaternary deposits.
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Figure 7.7: Back scattered electron micrographs in polished thin section.
Back scattered electron micrographs showing clinochlore (cli) and chlorite (chl) compatibility with various sulfide and carbonate minerals in drill core at ATO Pipe 1. A: Sample DDH ATO-02-49; sph, sphalerite; Q, quartz; smt, smithsonite. B: Sample ATO-02-146. Chlorite associated with pyrite (py) and galena (gn). C: Sample ATO-02-146. Same as B. H, hole in polished thin section. D: Sample ATO-02-146. Galena inclusions in pyrite and sphalerite mantled by chlorite (Figure 7.7).
Though quartz is the dominant alteration mineral at the surface, magnesium minerals also are widespread in the pipes. They, in essence, replace rock flour during pipe development and eventually comprise a fluidized matrix together with iron sulfide and base-metal sulfide minerals. Their importance is well indicated by mineralized core contents of many 10s of thousands ppm Mg to greater than 100,000 ppm Mg throughout the mineralized pipes. The dominant Mg alteration mineral in the pipes is clinochlore (hydrous Mg Al silicate member of the chlorite group). With increasing depth in the pipes, clinochlore becomes progressively enriched in Fe and assumes petrographic characteristics of typical chlorite and, in turn, becomes associated with increased abundances of phlogopitic white mica (Figure 7.7).
Figure 7.8: Aspects of various alteration assemblages typically present in drill core at ATO
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Typical appearance in core at ATO of the various altered rocks recognized thus include (1) oxidized siliceous sinter; (2) steeply dipping quartz veinlets with associated pyrite and base metals; (3) matrix supported mineralized breccia where magnesian chlorite is the dominant hydrosilicate in the matrix; and lastly (4) gypsum after anhydrite especially concentrated near pipe margins (Figure 7.8). As depicted in Figure 7.8 (ATO-35-37.75), many quartz veinlets containing abundant sulfide minerals can be steeply dipping in the pipes (i.e., essentially parallel to core axes). In addition, Mg chlorite (clinochlore) plus sulfide minerals also may be disseminated widely throughout heavily mineralized core where the rocks in effect are matrix supported by alteration silicates and sulfide minerals (ATO-38-95.85). These relations, together with presence of well-defined sulfide-silicate banding in many veins (see below), suggest veining in the pipes continued well after initial replacement of rock flour during the earliest stages of mineralization.
Because of the importance that the silica cap at Pipe 1 played in the discovery of the mineralized pipes at ATO, it warrants description in some detail.
7.3.3.1 Silica Cap in Pipe 1 Silica cap, though presently recrystallized multiple times during repeated passage of fluids streaming upwards through the pipe, has a number of characteristics that indicate it was originally deposited as colloform-banded sinter near the original paleosurface of an intermediate sulfidation system (see below). Silica cap rock is highly resistant and readily recognizable even from moderate distances because of its stark white color against a dark landscape (Figure 7.9).
Highly disrupted banded quartz veins comprise at silica cap rock in Pipe 1 at ATO. These silica outcrops are dominated by quartz, which, under the microscope, shows effects of repeated recrystallization, commonly marked by newly grown quartz transecting growth zones in previously crystallized colloform or banded quartz. Again, potassium feldspar is not present in these rocks.
Almost all silica cap rock encountered by drilling is oxidized and includes various abundances of secondary minerals and traces of primary sulfide minerals. Silica cap is made up predominantly of quartz (sparse chalcedonic fabrics are preserved) and variable concentrations of iron oxide minerals, as well as less abundant kaolinite, illite, numerous primary and a number of secondary Pb minerals (rare galena mostly preserved in a surrounding mantle of a Pb–Al phosphate (plumbogummite); pyromorphite (Pb phosphate); and Pb–Mn oxide minerals), sphalerite (rarely encapsulated in quartz), arsenopyrite (also in quartz), argentite, barite (in narrow micro veins as well as tabular crystals in open cavity fillings), clinochlore, chlorite, and rare prehnite.
It is the secondary Pb minerals at ATO that provide the source for the strong Pb anomaly in soils at ATO.
Outcrops on Pipe 1 indicate that, overall, its silica records a complex geologic history. The rocks display highly disturbed almost chaotic orientations even within individual outcrops of about 3-5 m wide. Banded and crustified silica has orientations that are extremely variable. Further, presence of now bladed silica that replaced earlier deposited calcite indicates that boiling had to have occurred near the top of the ATO system, wherein removal of CO2 after breakdown of bicarbonate led to deposition of early paragenetic stage, bladed calcite at the paleo uppermost levels of the silica cap. A boiling environment must have contributed to further disruption of the rocks, as well as a number of other geologic events. This boiling environment must have been below the water table underlying sinter at the actual paleosurface of Pipe 1.
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Figure 7.9: Photographs of outcrops of silica cap rock at Pipe 1 at ATO.
A: Ribs of silica marking margins of empty cavities formerly occupied by sulfide minerals, mostly pyrite but also a number of Pb minerals. B: Feeder vein of quartz (outlined in red) into basal part of banded quartz veins. C: Angular blocks of colloform-banded quartz veins engulfed by additional vein quartz. Note matchbox near center top of photo for scale. D: Close up view of recrystallized banded silica showing cavities formerly filled by mostly pyrite (Figure 7.9)
Some surface outcrops of silica cap initially must have crystallized almost at the Middle-Late Jurassic (?) paleosurface during mineralization at ATO. Presence of reticulated mats of silicified reeds both in longitudinal section and in cross section are well preserved in some grab samples (Figure 7.10). Most rock in the sample is quartz. Note oval and circular features present in Figure 7.10B. The dark material in both photomicrographs that outline the reeds is considered to be phosphorous-rich carbonaceous material relict from the reeds, which presumably were growing in thermal ponds at the actual paleosurface as mineralization at ATO was evolving at depth. Typically, phosphorous contents of drill core through silica cap are in excess of 1,500 ppm. We infer the fossil reeds at ATO to be somewhat analogous to reeds present today in thermal ponds surrounding the geysers at Yellowstone National Park. In fact, the micro plumose or feathery outlines of the reeds preserved in thin section at ATO are quite similar to those of present-day reeds at Yellowstone.
All introduced silica especially below (Figure 7.10) the surface shows complex recrystallization textures indicating repeated passage of fluids associated with base and precious metal introduction. Cross fiber textured quartz characteristic of chalcedony has not been found to date because of recurrent crystallization. However, mammillary-like growth zones are still locally preserved in some samples at the thin section scale suggestive of original deposition of silica as opaline silica. This especially pertains in the cap rock parts of the system. These growth zones are still visible through cross cutting, newly recrystallized quartz crystals. In light of the preserved relict reeds locally in now extinct thermal pools at the present day surface, the absence of widespread chalcedonic quartz and well-banded silica at depth suggests the pools may have been quite short lived. Undoubtedly, near-surface siliceous sinter and banded silica must have largely recrystallized as the sinter was broken and disrupted during repeated passage of mineralizing fluids. From the present day extent of the silica cap at Pipe 1, the paleo thermal field must not have had that wide a footprint.
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Figure 7.10: Relict reticulated mats of silicified reeds siliceous sinter. Plane polarized transmitted light.
Relict reticulated mats of silicified reeds in Middle-Late Jurassic (?) siliceous sinter preserved in grab samples of silica cap rock at Pipe 1, ATO deposit. A: Shows mostly longitudinal sections through reeds (at arrowhead). B: Mostly oval sections through reeds (at arrowheads) (Figure 7.10).
Most importantly, particles of free gold containing variable amounts of silver are present in surface samples of silica cap rock (Figure 7.11).
Presence of free gold in surface samples of silica cap at ATO was further verified using the SEM. Some of this free gold can have about a 60/40 ratio in its Au/Ag content, and gold has been shown by drilling to be especially concentrated throughout the silica cap portion of the underlying pipe. Potassium feldspar is exceedingly rare in these rocks, and is present only as < 1mm-wide grains, where it has an irregular wormy texture locally adjacent to iron oxide minerals.
Figure 7.11: Photomicrographs in reflected light of particles of free gold (Au) in surface grab samples
Figure 7.11 shows Photomicrographs in reflected light of particles of free gold (Au) in surface grab samples at ATO. Sample JB-AT-572 from silica cap at Pipe 1; Sample PS-TO-493 at Pipe 2.
Though recognition of Au in high-grade grab samples from surface outcrops proved to be relatively straightforward by both SEM and standard petrographic methods, this turned out to become increasingly difficult at depth as the mineralized system becomes highly enriched in galena. Because of the extremely close effective atomic numbers of Pb and Au, back-scattered imaging by SEM became an inefficient means of searching for Au. Widespread presence of galena in the samples made it impossible to test by EDAX all back-scattered bright areas for Au, because of the inordinate amount of time required. Nonetheless, some Au was found deep in the system during recon SEM study of core samples, and textural relations of Au in the three pipes are discussed more fully below.
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Additional photos showing textures of oxidized silica cap at ATO demonstrate how widespread and penetrating the former presence of sulfide minerals was at the top of Pipe 1 (Figure 7.12).
Figure 7.12: Photographs of silica cap rock in oxidized drill core from Pipe 1
Figure 7.12 showing variable concentrations of pre-oxidation sulfide minerals. A, C: Abundant iron oxide at margin of quartz vein. B: Silica ribs surrounding vacuoles partially filled by iron oxide similar to silica outcrop. D: Assay interval contains 1.44 ppm Au, 2.8 ppm Ag, 3,170 ppm Pb, and 296 ppm Zn.
7.3.3.2 Styles of Sulfide-mineralized Rock at ATO A number of styles of sulfide-mineralized rock are present below the oxide zone in the pipes at ATO, ranging from disseminated flooding by sulfide minerals in matrix of breccia to multiply banded veins. The latter may be either flat lying or steeply dipping. Generally, sphalerite becomes more Fe rich with depth in the system; compare honey brown sphalerite at 82.4 m (Figure 7.13A) with brown sphalerite at 97.80 m associated with weakly amethystine quartz (Figure 7.13B), or honey brown sphalerite at 85.9 m versus dark brown sphalerite at 142.7 m in DDH ATO-20 in Pipe #2 (Figure 7.14). The latter also is associated with amethystine quartz. Further, flat-lying galena-sphalerite flooding at 266.5 m in DDH ATO-111 can be followed down hole within a few meters by steep veins at 275.1 m (Figure 7.14). The latter veins also cut paragenetically earlier disseminated sulfide minerals that form a matrix support to mineralized breccia. In addition, late paragenetic stage manganiferous calcite, in places actual rhodocrocite, rarely cuts across locally layered breccia. In addition, the multiple bands of sulfide minerals and quartz in many veins suggest a protracted period of vein emplacement after initial onset of brecciation associated with pipe emplacement.
Overall, the mineralized system at ATO really does not contain that much manganiferous calcite, though some is present rarely in cm wide veins.
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Figure 7.13: General fabric of mineralized breccia in drill holes from Pipe 2, ATO.
Figure 7.13 shows general fabric of mineralized breccia in DDHs 19 and 20, Pipe 2, ATO deposit. A: Interval contains 1,783 ppm Cu and 33.9 ppm Ag. B: Note amythestine quartz associated with introduction of galena and sphalerite. Interval contains 21.1 ppm Ag, 47,200 ppm Pb, and 52,900 ppm Zn. C: Abundant calcite– impregnated breccia near pipe margin where breccia fragments are sugary textured, chilled margin of post- mineral dike. Interval contains 0.6 ppm Ag, 1,164 ppm Pb, and 2,196 ppm Zn. D: Note sphalerite (pale brown) and galena (blue gray). Interval contains 1.94 ppm Au, 5,738 ppm Cu, 146,900 ppm Pb, and 106,300 ppm Zn. E: Quartz-sphalerite banded veins. Interval contains 19.2 ppm Ag, 79,400 ppm Pb, and 133,400 ppm Zn. F: Breccia including angular fragments of pale brown metasiltite.
Figure 7.14: General styles of mineralized rock at ATO
General styles of mineralized rock at ATO (from left: ATO-20 – 85.9; ATO-20-142.7; ATO-111–266.5; ATO- 111-275.1) shown in Figure 7.14.
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7.3.4 Mineralization at ATO In summary with regards to mineralization at ATO:
- ATO is an intermediate sulfidation system (IS) Neutral low temperature, near paleo surface fluids, bladed silica after calcite indicates some local boiling. Related to Jurassic magmatic event Banded silica, broken sinter, repeatedly recrystallized at paleo top - ATO is confined to pipe bodies Multiple collapse and upward transport in the pipes, repeated brecciation followed by continued ingress of steep and shallow veins, veinlets and flooding - ATO is Mg chlorite and silica dominant system Quartz, clinochlore (high Mg, Al silicate), kaolinite, gypsum (peripheral after anhydrite), adularia is absent - ATO is dominated by free gold, base metal and Ag sulfides Lead-phosphates (near surface) – Pb-carbonates – galena (at depth) Zinc-carbonate (near surface) – low FeS sphalerite (at depth) Au in quartz & in sulfides; Ag mostly in tetrahedrite & miargyrite The mineralized pipes at ATO are unusual from a variety of standpoints. Nonetheless, a number of general statements can be made concerning their genesis and classification. As noted above, the ATO base (Pb-Zn- (Cu)) metal and Au-Ag mineralized pipes are intermediate sulfidation (IS) epithermal deposits characterized by an absence of adularia. They are apparently magma affiliated on the basis of their sulfur isotopic data, and must thus be associated with emplacement of Jurassic igneous rock at depth. The predominance of low FeS sphalerite, galena, Ag-bearing tetrahedrite-miargyrite, and chalcopyrite at ATO unquestionably are all compatible with an IS state (Figure 7.15). However, as opposed to a strong correspondence of IS deposits worldwide (Western US, Peru, central and Eastern Europe) with andesitic magmas (Sillitoe and Hedenquist, 2003), ATO appears instead to be associated with more mafic, dioritic Jurassic arc terrane magmas. These magmas thereby must have contributed to a correspondingly high Mg/2H+ component in fluids associated with mineralization that in turn inevitably led to widespread presence of the Mg chlorite clinochlore as a gangue mineral in the pipes as opposed to adularia. In addition, Jurassic volcanic rocks are not present in the region of ATO even though the pipes appear to have been formed near surface (banded silica, siliceous reed- bearing sinter at Pipe 1). This in turn suggests rapid (?) emplacement of the pipes occurred as a temporally transient or fleeting event. Regardless, well-formed sulfide-silicate banded veins in the pipes suggest protracted veining continued after formation of the brecciated columns of rock. However, the question still remains as to whether or not the mineralized pipes at ATO represent distal parts of a deep-seated porphyry environment.
Thus, upward flaring brecciation in Pipes 1, 2, and 4 at ATO provided fluid pathways for ingress of fluids associated with an open-space filling mineralization event that occurred over a protracted time interval. Mineralization in the pipes does not represent an instantaneous event essentially contemporaneous or immediately following the pipes’ piercing through their surrounding host rocks. Fist-sized veins, well banded with successive layers of early low FeS sphalerite and somewhat later galena, together with precious metals, in the pipes attest to a relatively protracted time for individual vein emplacement within the confines of the pipes.
Origin of brecciation remains enigmatic, however, and shows features of both magmatic and hydromagmatic breccias, as well as tectonic breccias (see Sillitoe, 1985). Additional complications hampering straight forward interpretation results from superposition of brecciation associated with mineralization onto detrital fragment-rich Early Permian volcaniclastic rock and Jurassic pebbly conglomerate, as well as protracted passage of fluids associated with mineralization in the pipes. Nonetheless, some fine-grained, matrix-
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT supported breccias with abundant rock flour encountered at depth at ATO NW of Pipe 2 are typical of magmatic breccia-style diatremes (magmatic hydrothermal systems that extend to surface). Rock flour where present at ATO, though important from a pipe genesis standpoint, is itself not mineralized, and represents those brecciated rocks that did not encounter mineralizing fluids and thus were spared alteration during subsequent passage of the fluids.
After near-surface deposition of banded siliceous sinter in an essentially horizontal orientation, sinter was disturbed into jumbled blocks as the process of underlying mineralization continued to evolve. This may account for the highly discordant attitudes of sinter layering among nearby outcrops at Pipe 1, though tectonic disturbance after pipe emplacement also may have contributed to the discordant attitudes. Banded and crustified silica has orientations that are extremely variable in outcrop at Pipe 1. Further, presence of now bladed silica that replaced earlier deposited calcite indicates that some boiling had to have occurred near the top of the ATO system, wherein removal of CO2 after breakdown of bicarbonate led to deposition of early paragenetic stage, bladed calcite at the paleo uppermost levels of the silica cap. A boiling environment also must have contributed to disruption of the rocks, as well as a number of other geologic events. This boiling environment must have been below the water table underlying the well-developed sinter at Pipe 1.
As described earlier in the report, the pipes at ATO were cut by a number of postmineral, fairly flat-lying faults and dikes. Norton and Cathles (1973) note that the primary question concerning any hypothesis of breccia development is how the void amongst the fragments was created. Voids in breccias typically provide sites that are filled partially to completely by subsequently introduced gangue (quartz, magnesian chlorite, phlogopitic white mica at ATO) and ore minerals (sphalerite, galena, chalcopyrite, pyrite, tetrahedrite /tennantite, miargyrite, and particles of free gold at ATO). As further noted by Sawkins and Sillitoe (1985), the largest variety worldwide of breccia types resides in magmatic arc terranes. Sillitoe (1985) lists the following six fundamental mechanisms for formation of breccias:
- Release of fluids from high-level magma chambers during second boiling; - Magmatic heating and expansion of meteoric pore fluids; - Cool ground waters interacting with magma; - Magmatic-hydrothermal brecciation, including pre- and postmineral diatremes; - Mechanical disruption of a magma’s wall rocks by subsurface movement; - Fault displacements. For a variety of reasons, at ATO the most likely mechanism associated with emplacement of Pipe 1, 2, and 4 is the fourth one listed above (i.e., magmatichydrothermal brecciation). Decompression of volatiles associated with second boiling in a dioritic complex at depth led to formation of the permeable channel ways that subsequently were filled by mineralized rock. Another characteristic of these types of breccias is that in many districts they tend to be present in clusters (Sillitoe, 1985), a characteristic that applies to ATO. Yet, what is observed at ATO and not at many other magmatic-hydrothermal breccia occurrences is upward pipe termination where at Pipe 1 now highly disturbed siliceous sinter (and highly Au-mineralized as well) is present very close to the original paleosurface.
Nonetheless, some differences from the norm for many well-described, mineralized magmatic-hydrothermal breccias elsewhere characterize the breccias at ATO. These differences include typically diffuse or poorly marked pipe margins at ATO attributable largely to a relatively prolonged mineralization event that continued well after cessation of the brecciation that first created the voids.
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Figure 7.15: Sulfur fugacity versus temperature diagram
Above sulfur fugacity versus temperature diagram showing various sulfide mineral assemblages in epithermal deposits that reflect sulfidation state varying from low through intermediate to very high sulfidation states. Approximate compositional fields of geothermal, magmatic hydrothermal, and volcanic fumaroles shown in varying shades of gray. Approximate position of higher temperature parts of ATO mineralized system shown in pink on the basis of sulfide mineral assemblages stable in pipes. From Sillitoe and Hedenquist (2003, Linkages between volcanotectonic settings, ore-fluid compositions, and epithermal precious-metal deposits, in Simmons, S.F., ed., Giggenbach Volume, Society of Economic Geologists and Geochemical Society, Special Publication 10) (Figure 7.15).
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8 DEPOSIT TYPE
8.1 Geological Model for ATO District
The geology of the property consists of metamorphosed Devonian sedimentary rock overlain by a volcanic and sedimentary sequence of Permian age and remnant scraps of probable Jurassic volcaniclastic units, intruded by Jurassic plutons ranging from diorite to granite in composition and including rhyolitic phases mainly as dikes. Petrographic study suggests simple, single-pulse injections of the intrusions, with latestage generation/expulsion of felsic phases and contemporaneous concentration of metal-rich fluids.
Mineralization on the property is probably related to the Jurassic magmatic rocks as sources of heat, metals, and fluids. The main ATO system is associated with a stratigraphic unit that appears to be localized in a graben or collapse feature that is possibly but not certainly Jurassic in age; any stratigraphy could potentially be prospective for this style of mineralization. The ATO mineralization appears to have occurred as a protracted single-stage process in separately upward-flaring pipe-shaped bodies, with temperatures ranging from ~300-3500C at depth to possibly ambient temperature in surficial sinter. The nearby Mungu mineralization also appears to have occurred in a single pulse, but at higher temperatures of ~375-4000C intimately associated with emplacement of rhyolite.
In a broad sense, metal zoning shows a clear, classic pattern of intrusion-centered copper (plus or minus molybdenum, tungsten, gold, and other elements) outward to country-rock hosted lead, zinc, and silver (plus or minus gold, arsenic, antimony, mercury, and other elements). It is presumed that this lateral zonation also occurs vertically, as evidenced by increasing copper values at depth in ATO.
The three main Jurassic intrusive phases are diorite-granodiorite, granite, and rhyolite. The diorite- granodiorite plutons are highly magnetic and typically develop large aureoles of magnetic hornfels. They show little or no quartz veining and exhibit no significant alteration apart from weak chloritization, which may be a regional metamorphic effect. On their upper and outer contacts they locally have patchy albite- sericite (muscovite) zones which are considered to be simple deuteric alteration related to final crystallization processes. In some cases they appear to be zoned inward to more felsic phases. Pegmatite- aplite phases show diffuse margins, suggesting segregation and streaming of more felsic fractions during late stage crystallization. Miarolytic cavities are common and locally abundant, containing coarse euhedral biotite, magnetite, pyrite, chalcopyrite, and local bornite.
The granite plutons are moderately magnetic and produce smaller, patchy hornfels aureoles. They may locally be zoned outward to more mafic composition. The upper and outer contacts typically show patchy to pervasive sericite (muscovite) alteration and common to ubiquitous gray quartz-sulfide veins which are often druzy. Pegmatite-aplite phases are common and locally show possible unidirectional solidification textures, and rhyolitic phases with diffuse margins are present, all suggesting segregation and streaming of volatile- rich phases during late-stage crystallization.
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Figure 8.1: Cartoon model of the Jurassic intrusives in the ATO district
Metal patterns and temperatures of mineralization associated with Jurassic intrusives shown in Figure 8.1 (green: diorite-granodiorite; pink: granite; tan: rhyolite; yellow: deuteric alteration; brown: hornfels; red: quartz; pink: sinter).
The rhyolitic phases are typically emplaced as dikes and small plugs. They are moderately magnetic but do not produce contact metamorphic aureoles. The rhyolites are typically pyritic, and locally highly so.
The metal patterns appear to vary systematically with intrusive composition. The diorite-granodiorite bodies have a copper-gold-tungsten signature related to visually obvious disseminated sulfide and sulfide-filled miarolytic cavities. The flanking hornfelsed country rocks generally show annular halos of geochemically anomalous lead and zinc, and locally contain percent-level concentrations of base metals plus or minus silver.
The granite bodies show essentially the same patterns, with some differences. The granite intrusions show copper anomalies, but typically at only geochemical levels, and the anomalies in the flanking country rocks have a more distinctly silver-rich character. The largest apparent difference, however, is the local development of a gold-bismuth-molybdenum zone in an intrusion-proximal setting within hornfels.
The rhyolites have different metal patterns depending on the level of emplacement. At deeper levels where the rhyolites were confined under lithostatic load, the geochemical signature is gold-silver-copper. In this setting it appears that sulfides may have been admixed with the rhyolite magma as it was being emplaced, and it is likely that metal deposition was caused by cooling and fluid mixing. At shallower levels under hydrostatic conditions the geochemical signature is lead-zinc-silver-gold, with metal deposition related to cooling, possible boiling, and possible wallrock interactions.
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There is relatively little hard data available on temperatures and isotopes for mineralization in the ATO district. Petrographic relationships would indicate that the disseminated and miarolytic cavity-filling sulfides in the diorite-granodiorite intrusives were deposited at magmatic temperatures of 400C or more, and the temperatures would be roughly equivalent or slightly lower for the granites. Mineral geothermometry using paired arsenopyrite and sphalerite in drill core from ATO gives ~275-3500C. The lower end of the possible temperature range is suggested by siliceous material at the outcrop of ATO Pipes 1 and 2, which has been interpreted as sinter based on textures and high phosphorus contents. 8.2 Genesis of the ATO Deposit
It is believed that the ATO deposit is an epithermal gold and polymetallic deposit of transitional sulfides in breccia pipes in a Mesozoic continental rift zone in eastern Mongolia. Below are accounts on similarity of ATO to other deposits in terms of formation of breccia pipe, its development stages, geochemical zonation, and the type of deposit genesis. 8.2.1 Formation of Breccia Pipe Tectonic-magmatic activities began to take place in the area around ATO deposit in Early Jurassic when hydrothermal and metasomatic alterations and mineralization were formed in relation to intrusions and dykes. Absolute age of an intrusive massive located outside and east of ATO prospect area was determined to 189 million years. In addition, an Early Jurassic massive was formed in the southern part of the area comprises two phases of rocks of sub-alkaline series distributed in small, separate outcrops of small intrusions. Criteria or signs for copper porphyry mineralization have been noticed in these two intrusions. It is believed that the pipe bodies of ATO deposit were formed in a structure that resulted from partial melting and upheaval of magmatic fluids. This can be explained in more detail in a model proposed by Noel White. In the area of ATO deposit, a small basin was deposited with sediments in Lower Jurassic, which include sandstone, siltstone, tuff sandstone, and gravelite, conglomerate, with coarser rocks at the bottom and finer rocks at the top. Quartz sinter was accumulated in pipe structures at ATO deposit as a result of hot spring activities (Figure 8.2A) at the same time as the formation of these sedimentary rocks. This can be explained by the facts that sinter materials and fragments of low temperature chalcedonic quartz vein are found in the gravelite, and that lenticular bodies of gravelite are found in the sinter.
Figure 8.2: Model of formation of ATO deposit
After the formation of layered quartz accumulation, or sinters, on the surface of the ground, the sinters were broken apart into blocks of varying sizes (Figure 8.2B). This explains the outcrop called Pipe 1. The layered quartz sinter at this outcrop is broken into blocks, and the layering in these blocks has become disordered and oriented in random directions. Sedimentary rocks in the pipe structures become brecciated allowing flow of fluids and providing a domain for mineralization.
Pipe-form mineralization were subjected to post-mineralization horizontal faulting and appear to be displaced along the fault (Figure 8.2C). Small bodies of diorite have been found in these faults.
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8.3 Geochemical Zonation
The pipes that have been identified in the deposit are aligned in a row in a structure with a trend from southeast to northwest and form mineralization with differing geochemical zonation with certain metallic concentrations. Noel White (2001) devised a pattern of flow of fluids based on this same geochemical zonation (Figure 8.3). According to him, it is possible that the intrusion that these mineralization were originated from may be located tens of kilometers away from ATO deposit.
Figure 8.3: Section Model of ATO deposit indicating relationship of the deposit with a copper porphyry intrusion
From the ore minerals, sphalerite is the most common mineral in the pipes of the deposit with light yellow sphalerite as the dominant variety. In addition, abundance of veins and veinlets of gypsum indicate the oxidizing environment of the fluids and original magma. It is very likely that such kind of magma may be connected to porphyry mineralization (Figure 8.3 [1]). Therefore, there might be present an intrusion with copper porphyry (Figure 8.3 [1]), surrounding gold and copper molybdenum stockwork (Figure 8.3 [2]), and high sulfidation epithermal mineralization on top of them (Figure 8.3 [3]). Magmatic fluids with high metallic content (Figure 8.3 [5]) rise upward, halt at certain level, and drift horizontally along open spaces to get mixed with underground water and change its composition. After that, they break the burdens at some points to travel out to the surface through several different routes (Figure 8.3 [6]). This is how the pipe bodies of ATO were formed. Concentration of metallic components varies from pipe to pipe due to differences between pH levels, variations in pressure and temperature, and the differing composition of the fluids that were injected to the pipes of ATO deposit. It ranges from copper rich to lead and zinc rich and this explains the geochemical zonation in the deposit. As for ATO deposit, the burden that kept the fluids from traveling upward is thought to be a unit of very dense and massive, black colored siltstone. The siltstone unit occurs at the bottom of some of the deeper drill holes having thickness ranging from a dozen meters to 30-40 meters. A break in the black siltstone unit caused hydrothermal explosion and rapid upheaval of CO2, which resulted in formation of breccia pipes. Fluids rose through open spaces between breccia fragments where pressure and temperatures were lower and gangue and ore minerals filled them completely or partially. As a result mineralization took place. 8.4 Deposit Type
The ATO deposit is classified as an epithermal deposit with intermediate sulfidation based on the mineralization texture, geochemical associations, and mineral composition. The classification process involved comparison of the ATO deposit with natures of the high, intermediate and low sulfidations of epithermal deposits as shown in Figure 8.4.
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Figure 8.4: Nature and types of epithermal mineralization
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9 EXPLORATION
9.1 Preparation Stage
At this stage, reports and related materials, including maps, sketches, and logs, of historical mapping and prospecting works conducted by earlier researchers were collected, compiled and studied. Interpretation of aerial and satellite images were also performed. Coordinates used by CGM for exploration on the project is UTM coordinates with the datum set to WGS-84, Zone 49N. The boundary coordinates of the exploration and mining licenses are defined by latitude and longitude coordinates. 9.2 Field Work
The field investigation that was conducted in the license area can generally be subdivided into two stages: prospecting and exploration.
In 2003-2009, geologists of COGEGOBI, who had been specialized in prospecting and exploration of uranium projects, conducted geological mapping and prospecting traverses and collected geochemical samples in the exploration license area, supplemented by a magnetic survey. The result was they discovered an epithermal gold occurrence.
In 2010, CGM carried out a prospecting stage consisting of geological mapping and prospecting traverses, surface and other sampling tasks, variety of geophysical surveys, and some trenching and drilling. As a result of the prospecting ATO occurrence was chosen to a detailed study, and an intensive drilling program began in late 2010 to advance the project to exploration stage. 9.2.1 Geological Mapping and Prospecting Traverses Systematic mapping and prospecting traverses carried out in the entire ATO exploration license area. Traverses were placed 100 m apart. During traverses, grab samples were collected from alteration zones and rocks with possible mineralization. As a result, a 1:25,000 scale geological map was produced covering the entire license area.
Total of 397 grab samples were collected in 2010 during mapping. Systematic mapping continued until 2014 including detail mapping at the ATO prospect, as well as recon scale mapping with some grab samples.
Mapping and prospecting work was assisted by maps such as 1:32,000 scale aerial image, a satellite image, and topographic maps at scales of 1:25,000 and 1:50,000.
As a result of the work described above, the area contains the ATO gold-polymetallic occurrence chosen as a potential target. Detailed prospecting work was aimed to draw a shape and size of a mineralized body, boundaries of lithology and alteration zones, and study the sources of secondary dispersion halos and geophysical anomalies. Based on the result of the detailed prospecting traverses, detailed 1:10,000 scale geological map of the ATO deposit area was interpreted. 9.2.2 Field Sampling for Geochemical Analysis 9.2.2.1 Stream Sediment Sampling The ATO district is characterized by limited and weak systems of streams and galleys, wherein down stream sediment transport is practically non-existent. Thus, minimal stream-sediment sampling was carried out to obtain a general geochemical characteristic of the area in 2010.
As an initial task of the field work, stream sediment samples were taken from Holocene ditches and ravines, where outcrops of Late Paleozoic to Early Mesozoic rocks distributed, to detect stream sediment anomalies.
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Samples were taken using hand augers and shovels to dig 0.2-0.6 m deep holes and extract 1.5-2 kg samples from alluvial gravel and sand materials. To avoid contaminating the samples, stainless steel drills and sifters were used. Samples were air dried before sieved in a 20 mesh sifter. The samples were reduced to 150-200 gr after sieving and the reduced samples were submitted to the laboratory for analyses. A total of 509 stream sediment samples were collected counting 1 sample for per square km. Coordinates of sample location were recorded by a field GPS devices, and a simple field illustration was drawn including catchments width, depth, rank, and direction; and the sand ratio, gravel and clay size, color, degree of gravel rounding of samples; and information on nearby country rocks. A stream sediment map was produced as a result and used for further studies.
9.2.2.2 Soil Sampling Initial soil sampling (20 by 100 m) was completed at the ATO prospect in the fall of 2009 by CGM. During 2010, an additional 4,256 samples were collected at 100 by 100 m and 200 by 200 m in areas showing potential, and, during 2012, an additional 18,471 soil samples were collected from broad areas across the entire ATO district (Figure 9.1). These 18,471 soil samples are from seven domains within the ATO district, and include 7,290 samples collected from the Davkhar Tolgoi area south of the mineralized pipes at ATO, and 4,860 soil samples are from areas close to the pipes. Thus, broad expanses of the licensed areas to the NNE of the ATO deposit finally were covered in 2012 by systematically gathered soil grids, areas of the seven exploration licenses held by CGM at the end of 2012 that previously had not been sampled.
The best indicator of presence of significant, underlying gold-mineralized rock at ATO was a gold anomaly in soil at ATO. This gold anomaly in soil at ATO was as much as 600 m wide in its longest dimension and includes concentrations of 50–500 ppb Au. By far, this is the strongest gold anomaly in the immediate area of the pipes within ATO district.
In addition, a strong lead anomaly in soil is mostly coincident with the soil gold anomaly and contains generally 50–300 ppm Pb. Most lead in soil is derived from secondary lead minerals, including plumbogummite (a Pb-Al phosphate), pyromorphite (a Pb phosphate), and Pb–Mn oxide minerals in the uppermost, oxidized parts of the underlying mineralized pipes. In addition, anomalous lead in soil E–SE of the Bayangol prospect appears to be defined, in part, by the trace of an inward-dipping low-angle thrust fault exposed in some trenches, though Early Jurassic intrusive rocks emplaced into Early Permian volcanic rock, largely rhyolite, also may be contributing to the lead anomaly.
Further, anomalous >500 ppm Zn in soil is present in soil on top of Quaternary-covered mineralized rock in Pipe 4, as well as on the northern fringe of the surface projection of Pipe 1 and on the SW margin of Pipe 2.
Extensive soil sampling was undertaken in 2013 and a minor amount of soil sampling was undertaken in 2014 to infill the grid spacing in various areas on the property, and to slightly expand the grids beyond their previous coverage in a few places. Most areas were infilled from a pre-existing 200 m X 200 m grid spacing to 100 m X 100 m and 100 m x 50 m and 50 m x 50 m on target areas. The infill sampling enhanced some anomalies detected in the wider spaced grids, but did not identify any new significant anomalies.
Occurrence of rock outcrops was sparse in the license area; most of it was covered by loose sediments. Soil samples were extracted from a depth of 30-60 cm (B horizon), or form underneath the brown soil with plant roots, sometimes from 1 to 2 meter depths drilled by hand augers when the cover was significantly thick. Each weighing 1.5-2 kg, the samples were dried and sieved in 80 mesh steel sifters to produce 150-200 gr samples ready for analyses. Coordinates of sampling sites were recorded in GPS devices, and a journal was kept about the composition, color, and structure of soil, the depth of extraction, and information on nearby rock bodies.
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Figure 9.1: Map showing soil sample locations on gold patterns in ATO District during 2010-2014 (CGM 2014ER).
Red boundary illustrates MV-17111 license area (previous boundary) 9.2.3 Other Sampling tasks To evaluate the potentiality of occurrences, mineralized points and altered rocks identified by the prospecting and exploration work, and to determine elemental grades, size and shape, and boundaries of mineralization, samples such as grab samples, channel samples were collected.
9.2.3.1 Grab sampling Grab samples were collected from outcrops and fragments of altered and/or mineralized rocks during prospecting and mapping traverses in order to determine and evaluate the grades of valuable elements. Sometimes a composite sample was taken with certain intervals from sites where there extensive alteration and mineralization. A grab sample weighed 1-2 kg. Each of the samples was placed in a proper bag and its coordinate was recorded. A total of 422 such samples were collected during the program.
9.2.3.2 Channel samples After making a description and documentation of a trench, the trench was cleaned and a channel sample was taken from the walls and floor of the trench using a chisel and a sledge hammer. Mineralized bodies and
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT alteration zones found in trenches were sampled over 1 to 2 m. Unaltered rocks were sampled with intervals up to 5 m. With cross-section measured 10 by 5 cm, the samples weighed 8 to 15 kg. Locations of the samples were recorded using a differential GPS device. A total of 7,689 channel samples were taken.
9.2.3.3 Sampling for metallurgical tests A minimum of 500 g sample from each of selected diamond drill holes was submitted to the laboratory of Actlabs Asia for Bottle-roll test for gold. When selecting samples, they were sorted with regard to their grades of gold and Pb-Zn, degree of oxidation (oxidized, intermediate, and unoxidized), and the type of host rock, and they were selected for their relative consistency of distribution and ability to represent their respective mineralized bodies (Table 9.1). Sampling was aimed to test the metal recovery of mineralized bodies. A total of 93 metallurgical samples were prepared.
Table 9.1: Schedule of metallurgical samples
Grade Very high High Medium Low
Degree of
oxidation zed
Oxidized Oxidi Oxidized Oxidized
Transition Transition Transition Transition
Unoxidized Unoxidized Unoxidized Unoxidized Au + + + + + + + + + + + + Au-Pb-Zn + + + + + + + + + + + + Pb-Zn + + + + + + + + + + + +
In addition, a contract was established with the laboratory of Xstrata Process Support Center to perform metallurgical test, and pursuant to it, initial test samples were sent to the laboratory in April 2011 followed by the second test samples in July 2011. The testwork involved using of gravity method to produce low-grade concentrates and then flotation method for separating lead and zinc. The purpose of the testwork was to maximize metal recovery from the ore and solving the issue of process plant design. Below are accounts on the two stages of testwork.
1. The initial stage of the testwork was performed on five sets of samples representing oxidized, intermediate, and unoxidized zones of the upper and lower parts Pipe 1 and 2 (Table 9.2). Eleven drill holes were selected and the samples from them were sorted according to their high, medium or grades of gold and lead-zinc.
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Table 9.2: Sets of samples for initial metallurgical test
Zone Oxidation Drill Hole ID Intervals (m) Width (m) Set No Weight (kg) Upper Oxidized ATO-12 2.00 44.30 45.6 ATO - 1 128.8 ATO-14 2.60 27.30 14.0 ATO-15 0.90 13.00 4.0 Transition ATO-07 32.70 54.90 22.2 ATO - 2 127.7 ATO-11 38.90 75.90 37.0 ATO-14 66.95 78.05 11.1 Unoxidized ATO-11 75.90 97.00 21.1 ATO - 3 127.8 ATO-11 101.40 120.20 18.8 ATO-27 74.35 88.35 14.0 ATO-28 61.15 74.65 13.5 Lower Unoxidized (Fresh) ATO-15 108.70 154.25 45.6 ATO - 4 127.3 ATO-32 140.20 154.20 14.0 ATO-38 128.05 132.05 4.0 ATO-07 147.40 150.55 3.2 ATO - 5 128.8 ATO-07 162.30 170.10 7.8 ATO-12 125.00 127.00 2.0 ATO-12 130.00 136.00 6.0 ATO-14 94.10 100.35 6.3 ATO-24 124.40 144.40 20.0 ATO-34 189.50 209.20 19.7 Total 640.4
2. The second stage of testwork was performed on nine sets of samples. First seven of the nine sets of samples were taken from the following locations, respectively (Table 19). Weakly mineralized intervals with the minimum grades of Au and Pb-Zn. Intervals with certain grades but not included in resource blocks Intervals with extreme gold grades and moderate Pb-Zn grades Intervals with high gold grades and high to moderate Pb-Zn grades Intervals with moderate gold grades and extreme Pb-Zn grades Intervals with weak gold grades and moderate to weak Pb-Zn grades Intervals with mixed grades of gold and Pb-Zn
In addition, one of the two remaining sets was chosen in order to evaluate the metal recovery of the newly discovered Pipe 4. The other set was taken with an aim to increase the metal recovery of Pipe 2.
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Table 9.3: Sets of samples for second metallurgical test
Drill Hole ID Intervals (m) Width (m) Set No Weight (kg) Weakly mineralized ATO-128 70.00 82.00 Set-1 75.98 ATO-162 50.70 62.60 Outside of pipes ATO-55 114.00 122.00 Set-2 74.58 ATO-160 30.00 45.80 Mixed-1 ATO-64 115.80 121.80 Set-3 36.00 ATO-92 108.30 111.30 Mixed-2 ATO-41 60.10 69.10 Set-4 30.88 Mixed-3 ATO-60 110.95 119.10 Set-5 35.61 Mixed-4 ATO-60 81.05 89.40 Set-6 31.72 Master ATO-19 85.70 91.70 Set-7 249.95 ATO-40 50.25 58.40 ATO-49 101.00 106.00 ATO-71 74.60 84.60 ATO-87 172.10 177.30 ATO-110 55.30 62.55 ATO-135 50.50 69.00 ATO-136 160.90 174.90 Pipe 4 ATO-96 109.80 140.00 Set-8 200.29 ATO-116 25.00 50.00 Pipe 2 ATO-20 60.10 88.40 Set-9 168.27 ATO-55 135.05 145.50 ATO-60 119.10 141.80 Total 9.00 903.29
To prepare these samples, all of the half cores from the selected drill holes were retrieved from storage. Each of the sets was packed in a 60 L barrels, and all necessary documents were obtained for customs clearance. The samples were shipped via freight forwarder DHL. The samples weighed a total of 1543.7 kg.
9.2.3.4 Samples for physical characteristics Geotechnical samples were selected to have consistent distribution in the pipes of the deposit taking into account the types of rocks present, degree of oxidation of pies, and grades of gold and other metals. The samples were prepared in two different forms before submitting them to the Actlabs laboratory (ACTLABS).
10 cm long sample accounting for ¼ of the core sample. 117 such samples were tested for bulk density by coating them with paraffin and dipping them in distilled water to measure displaced water. The volume obtained was then compared with density of the distilled water to determine the bulk density of rocks. 124 samples each weighing 50 g were taken from remnants of the pulverized samples that had been prepared for assays. The pulverized samples were put into fluids in a standard condition, the volume of displaced water was measured, and it was compared with the density of the distilled water to determine the ultimate relative density.
9.2.3.5 Samples for Petrographic analyses To study the lithological composition and mineral composition of rocks found in the deposit area, a total of 45 samples were taken from all types of rocks and alterations. Their petrographic descriptions have been made by Professor Bal-Ulzii (PhD) of University of Science and Technology of Mongolia and Ph.D A.B. Ted Theodore of USGS, an adviser to CGM. Ted Theodore used in his petrographic analyses Zeiss Axioskop 40 Pol
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT microscope with zooms at 2.5X, 5X, 10, 20X, and 50X and a lens 10X-20 Pol capable of zooming at 500X. The microscope is able to work with reflected and absorbed lights. The microscope was equipped with a digital camera MicroPublisher 3.3 RTV, which was connected to an iMac computer with 2.4 GHz Intel Core 2 Duo processor. Photos taken with this camera was then processed in Adobe Illustrator CS5 and Adobe Photoshop CS5 to jpeg format. The results were presented in the previous sections describing ATO deposit styles and geological settings.
9.2.3.6 Mineralogical samples To study the mineral composition of the deposit’s mineralization, possible sequence of concentrating of minerals, and the nature of structure and texture of the mineralization, a total of 59 samples were taken representing all of mineralized assemblages at the deposit, and they were analyzed by senior teacher Myagmarsuren of the University of Science and Technology and Ph.D A.B. Ted Theodore of USGS, an adviser to CGM. Ted Theodore conducted his analyses in Menlo Park of USGS in California. He used a LEO 982 electronic microscope to determine the sequence of crystallization of minerals and their textures. He also determined chemical compositions at some random points in the samples. Necessary photos were taken and have been included in the report in jpeg format and results are presented in previous sections.
9.2.3.7 Samples for determination of absolute ages In order to determine the ages of extrusive and intrusive rocks mapped in the area, seven samples were taken and sent for analyses by U-Pb dating on zircon crystals to Ph.D. J.K. Mortensen of the Pacific Centre for Isotopic and Geochemical Research of the University of British Columbia.
15 to 20 zircon crystals were picked from each of the samples were analyzed by the Laser Ablation ICP-MS (described by Tafti, 2009) using a New Wave UP-213 laser. Zircon crystals that measure no less than 74 microns were selected and mounted in an epoxy puck along with several crystals of internationally accepted standard zircon (Plesovice, FC1) that dates 197 million years and a couple of internal quality control samples, and they were fed into the instrument in a linear form. Crystals selected were washed by reduced nitric acid for about 10 minutes and then rinsed in distilled water to make them high quality crystals with no alterations and no prints. Laser beam level was taken at 45% to allow ablation crater size to be 15 microns. Laser beam was off for 10 seconds and of for 35 seconds. Data collected was reduced in GLITTER software. Biases were corrected using the Plesovice standard. Adjustment of the instrument was controlled by zircons of the internal quality control. The analytical regime included 4 measurements of Plesovice’s standard zircon, two measurements of internal quality control, and 5 measurements of prepared zircons. Interpretation of the processed results was performed using ISOPLOT software of Ludwig.
9.2.3.8 Paleontological samples Six samples were taken from the paleontological faunal and floral relics found in the exploration area and they were analyzed to determine stratigraphic ages. The task was performed by PhD Minjin of the University of Science and Technology of Mongolia, who is a member of the Stratigraphic Commission of Mongolia. He delivered his descriptions of the samples along with photographs. 9.3 Exploration Results
9.3.1 Geochemical Analysis Results of the soil sampling at the ATO license area were produced to analyze dispersion halos for elements of Au, Ag, Pb, and Zn. Analysis was attempted to determine the vertical and horizontal zonations of elements associated geochemically to the ATO deposit and figure out the geochemical features of the deposit.
In order to determine the horizontal zonation of the deposit, 959 samples were analyzed (analyzed for a suite of 49 elements by ICP in the Stuart Global Laboratory) that belong only to the ATO deposit area. In Page 59 of 209
NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT addition, certain representative drill holes were selected at each of the pipes and their samples were analyzed for a suite of 45 elements by IPC in the ACTLABS laboratory. A total of 2651 core samples from a total of 19 drill holes were included in the study.
Elemental associations differ with verticality and horizontality for the mineralized pipes of the ATO deposit and their grades vary.
9.3.1.1 Horizontal geochemical zonation of the deposit At the ATO deposit, elements such as Au, Ag, Pb and Zn have spatially coincident anomalous grades. Therefore, here, anomalous grades of these elements are combined to make an integrated dispersion halo, which was shown in Figure 9.2. The background grades of elements are 8.7 ppb Au, 0.44 ppm Ag, 21.7 ppm Cu, 23.6 ppm Pb, and 63.5 ppm Zn.
Pipe 2
Pipe 1
Pipe 4
Figure 9.2: Combined dispersion halo at the ATO deposit
A geochemical anomalous halo of combined elements is outlined with a size of 750 by 900 m, which is slightly elongated to north. The pipe-shaped mineralization is coincident with and clearly distinguished by the contour of combined value of 30. This contour is generally irregular in shape. Pipe 1 has the highest concentration of metals in soil whereas Pipe 2 has low intensity. Anomalous contours have gradual weakening to the southwest and south of Pipe 1, which reflects the transportation of elements along the features of relief. Very high intensity is attained over a small distance at the northeastern part of the pipe. Degree of correlation between any two of important elements at the ATO deposit can be seen for each of the pipes in the tables below.
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Table 9.4: Correlations of elements in geochemical dispersion halo of Pipe 1
Correlation Au Ag As Cu Mn Mo Pb Sb Zn Au 1.00 0.66 0.53 0.44 -0.40 0.14 0.46 0.52 -0.04 Ag 0.68 1.00 0.56 0.60 -0.41 0.44 0.59 0.54 0.01 As 0.53 0.56 1.00 0.85 -0.33 0.56 0.76 0.82 0.30 Cu 0.43 0.60 0.85 1.00 -0.44 0.48 0.92 0.64 0.15 Mn -0.40 -0.41 -0.33 -0.44 1.00 -0.33 -0.30 -0.31 0.12 Mo 0.14 0.44 0.56 0.48 -0.33 1.00 0.37 0.49 0.23 Pb 0.46 0.59 0.76 0.92 -0.30 0.37 1.00 0.54 0.06 Sb 0.52 0.54 0.82 0.64 -0.31 0.49 0.54 1.00 0.06 Zn -0.04 -0.01 0.30 0.15 0.12 0.23 0.06 0.06 1.00 Based on analyses of 62 samples
Table 9.5: Correlations of elements in geochemical dispersion halo of Pipe 2
Correlation Au Ag As Cu Mn Mo Pb Sb Zn Au 1.00 0.63 0.46 0.76 -0.14 0.07 0.73 -0.14 0.08 Ag 0.63 1.00 0.59 0.50 -0.09 0.38 0.51 -0.02 0.12 As 0.46 0.59 1.00 0.39 -0.06 0.60 0.33 0.34 0.35 Cu 0.76 0.50 0.39 1.00 -0.16 -0.01 0.64 -0.08 0.33 Mn -0.14 0.09 0.06 -0.16 1.00 -0.04 0.09 0.19 0.54 Mo 0.07 0.38 0.60 -0.01 0.04 1.00 0.05 0.48 0.07 Pb 0.73 0.51 0.33 0.64 0.09 0.05 1.00 -0.06 0.31 Sb -0.14 0.02 0.34 -0.08 0.19 0.48 0.06 1.00 0.09 Zn 0.08 0.12 0.35 0.33 0.54 0.07 0.31 0.09 1.00 Based on analyses of 46 samples
Table 9.6: Correlations of elements in geochemical dispersion halo of Pipe 4
Correlation Au Ag As Cu Mn Mo Pb Sb Zn Au 1.00 0.59 0.88 0.81 -0.42 0.68 0.86 0.61 0.71 Ag 0.59 1.00 0.79 0.82 -0.08 0.84 0.72 0.89 0.89 As 0.88 0.79 1.00 0.97 -0.19 0.88 0.98 0.81 0.89 Cu 0.81 0.82 0.97 1.00 -0.08 0.94 0.97 0.88 0.95 Mn -0.42 -0.08 -0.19 -0.08 1.00 0.11 -0.20 0.05 0.00 Mo 0.68 0.84 0.88 0.94 0.11 1.00 0.84 0.90 0.96 Pb 0.86 0.72 0.98 0.97 -0.20 0.84 1.00 0.77 0.86 Sb 0.61 0.89 0.81 0.88 0.05 0.90 0.77 1.00 0.95 Zn 0.71 0.89 0.89 0.95 0.00 0.96 0.86 0.95 1.00 Based on analyses of 25 samples
Gold in Pipe 1 has moderate correlations with silver, arsenic, antimony, and lead, and a weak correlation with copper, which in turn has a very good correlation with lead. At Pipe 2, gold gives moderate correlations to copper, lead, silver, and arsenic. At Pipe 4, it has high correlations with lead, zinc, and copper, and moderate correlations with arsenic, antimony, and molybdenum.
9.3.1.2 Vertical geochemical zonation at the deposit In order to investigate the deposit to depth, some drilling sections were selected to represent the general characteristics and features of the deposit and used the analytical data of core samples from the mineralized pipes to establish the general pattern of vertical geochemical zonation by converting the intervals of core samples to elevation levels and comparing the grades of gold and associated polymetallic elements.
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Pipe 1
A total of 1329 samples from eight drill holes were assayed for 45 elements for Pipe 1. However, the table below shows correlations of only the following elements, which have been selected because of their importance in hydrothermal explosion deposits (Table 9.7).
Table 9.7: Correlations of elements in core samples from Pipe 1
Correlation Au Ag As Cd Cu Mo Pb Sb Zn Au 1.00 0.37 0.16 0.38 0.46 -0.05 0.52 0.05 0.38 Ag 0.37 1.00 0.47 0.49 0.56 0.28 0.57 0.34 0.49 As 0.16 0.47 1.00 0.08 0.23 0.33 0.18 0.46 0.08 Cd 0.38 0.49 0.08 1.00 0.68 0.08 0.78 0.25 0.98 Cu 0.46 0.56 0.23 0.68 1.00 0.04 0.68 0.43 0.70 Mo -0.05 0.28 0.33 0.08 0.04 1.00 0.02 0.17 0.07 Pb 0.52 0.57 0.18 0.78 0.68 0.02 1.00 0.22 0.77 Sb 0.05 0.34 0.46 0.25 0.43 0.17 0.22 1.00 0.27 Zn 0.38 0.49 0.08 0.98 0.70 0.07 0.77 0.27 1.00
Gold in Pipe 1 has moderate correlations with silver, cadmium, copper, and zinc, and copper has strong correlations with cadmium, lead, and zinc, and moderate correlations with silver and antimony. Lead has strong correlations with cadmium, copper, and zinc, a moderate correlation with silver, and a weak correlation with antimony. Zinc has a very strong correlation with cadmium, strong correlations with copper and lead, a moderate correlation with silver, and a weak correlation with antimony.
Figure 9.3: Gold and other elements in Pipe 1 and their correlation at depth
Note: Vertical axis represent grades of elements in ppm and the horizontal axis represents depth in meters.
A depth analysis for correlation changes of the gold mineralization in Pipe 1 continues from the 880 m level to the 1070 m level. As per polymetallic elements, their mineralization occurs at two separate depth ranges such as from level 810 m to level 860 m and from level 880 m to level 1070 m. The lower range of polymetallic Page 62 of 209
NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT mineralization has no gold and discontinuous while the mineralization at upper levels coincides with gold mineralization and feature discontinuous high grades. Grades of cadmium, zinc, and copper decrease dramatically near surface, or at the 1060 m level, while those of gold, silver, and lead remain relatively stable.
Pipe 2
A total of 374 samples from three drill holes were assayed by multi-element analyses for Pipe 2. The results were then processed and displayed in Table 9.8 showing correlations of gold and polymetallic elements.
Table 9.8: Correlations of elements in core samples from Pipe 2
Correlation Au Ag As Bi Cd Cu Mo Pb Sb Sn Zn Au 1.00 0.28 -0.01 0.11 0.43 0.38 -0.03 0.18 0.08 0.30 0.43 Ag 0.28 1.00 0.18 0.25 0.79 0.87 0.15 0.96 0.66 0.72 0.79 As -0.01 0.18 1.00 -0.02 0.07 0.05 0.09 0.16 0.37 0.03 0.06 Bi 0.11 0.25 -0.02 1.00 0.29 0.24 0.00 0.25 0.16 0.32 0.29 Cd 0.43 0.79 0.07 0.29 1.00 0.86 0.09 0.73 0.47 0.87 1.00 Cu 0.38 0.87 0.05 0.24 0.86 1.00 0.13 0.79 0.59 0.79 0.86 Mo -0.03 0.15 0.09 0.00 0.09 0.13 1.00 0.14 0.22 0.15 0.09 Pb 0.18 0.96 0.16 0.25 0.73 0.79 0.14 1.00 0.63 0.66 0.72 Sb 0.08 0.66 0.37 0.16 0.47 0.59 0.22 0.63 1.00 0.44 0.46 Sn 0.30 0.72 0.03 0.32 0.87 0.79 0.15 0.66 0.44 1.00 0.88 Zn 0.43 0.79 0.06 0.29 1.00 0.86 0.09 0.72 0.46 0.88 1.00
Gold has moderate correlations with cadmium, copper, tin, and zinc, and weak correlations with silver, bismuth, and lead. Copper has strong correlations with silver, cadmium, and zinc, a moderate correlation with antimony, and weak correlations with bismuth and molybdenum. Lead has a very strong correlation with silver, strong correlations with cadmium, copper, antimony, tin, and zinc, and weak correlations with arsenic, bismuth, and molybdenum. Zinc has a perfect correlation with cadmium, strong correlation with silver, copper, lead, and tin, a moderate correlation with antimony, and a weak correlation with bismuth.
Figure 9.4: Gold and other elements in Pipe 2 and their correlation at depth
Note: Vertical axis represent grades of elements in ppm and the horizontal axis represents depth in meters.
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According to the depth analysis in Pipe 2 (Figure 9.4), gold occurs at a zone from 910 m level to 930 m level with very high grades and at a zone from 975 m level to 1050 m level with irregular lower grades. As for other metals, there is a zone of irregular high grades from 910 m level to 960 m level, followed by a zone of lower grades from 960 m level to 1050 m level. The graphs show that the grades of copper, silver, and lead increase slightly near surface whereas grades of zinc and cadmium start to drop at 1040 m level to very low grades.
Pipe 4
Correlations of elements in Pipe 4 are provided in the table below.
Table 9.9: Correlations of elements in core samples from Pipe 4
Correlation Au Ag As Cd Cu Mo Pb Sb Zn Au 1.00 0.11 0.06 0.26 0.36 0.03 0.19 0.12 0.23 Ag 0.11 1.00 0.05 0.07 0.11 0.17 0.09 0.06 0.09 As 0.06 0.05 1.00 -0.02 0.07 0.40 0.04 0.50 0.01 Bi 0.01 0.00 0.00 0.04 0.02 -0.01 0.04 -0.01 0.03 Cd 0.26 0.07 -0.02 1.00 0.75 0.07 0.63 0.33 0.97 Co -0.12 -0.08 -0.05 -0.15 -0.17 -0.05 -0.11 -0.12 -0.16 Cu 0.36 0.11 0.07 0.75 1.00 0.10 0.61 0.49 0.77 Mo 0.03 0.17 0.40 0.07 0.10 1.00 0.17 0.37 0.09 Pb 0.19 0.09 0.04 0.63 0.61 0.17 1.00 0.33 0.67 Sb 0.12 0.06 0.50 0.33 0.49 0.37 0.33 1.00 0.36 Zn 0.23 0.09 0.01 0.97 0.77 0.09 0.67 0.36 1.00
In Pipe 4, gold has a moderate correlation with copper and weak correlations with silver, cadmium, lead, antimony, and zinc. Copper has strong correlations with cadmium, lead, and zinc, a moderate correlation with antimony, and weak correlations with silver and molybdenum. Zinc has very a strong correlation with cadmium, strong correlations with copper and lead, and a moderate correlation with antimony. These suggest that polymetallic elements have better correlations with each other than with gold for Pipe 4. This is also seen clearly on the graphs of grades of gold and polymetallic elements displayed in Figure 9.5.
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Figure 9.5: Gold and other elements in Pipe 4 and their correlation at depth.
Note: Vertical axis represent grades of elements in ppm and the horizontal axis represents depth in meters.
Elements such as copper, lead, cadmium, and zinc have irregular pattern of grades at depth ranges of 860- 900 m level, 910-925 m level and 925-1050 m level. As per gold, its mineralization is established at depth ranges of 860-885 m level and 910-1050 m level. In general, gold and polymetallic mineralization has been identified having stable and relatively thick distribution below surface. This near-surface zone is marked by good correlations of gold, silver, cadmium, copper, lead and zinc, while the zones of high grades of certain metals with irregular distribution and low thickness that are found below it do not demonstrate correlations with gold; only the polymetallic elements have good correlations with one another. 9.3.2 Trenching A significant trenching program was carried out in 2010 and 2011 to test geochemical anomalies and the geological environment mainly in the ATO prospect but also in the surrounding Davkhar Tolgoi, Bayan Munkh, Bayan Gol and Duut Nuur occurrences. In some cases, trenching led to new target discoveries. Trenching was concentrated mainly in ATO deposit area from 2012 to 2014. In all, 244 trenches were excavated in ATO district surrounding areas including 168 trenches in ATO prospect were conducted totaling of 28,809 m work done.
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Table 9.10: Significant trenching results
Length Au As Zn Trench ID From (m) To (m) Lithology Ag ppm Cu ppm Pb ppm (m) ppm ppm ppm ATO-TR-117 2.2 2.8 5.0 Tuff dacite 0.13 < 201 11 15 89 1.2 19.5 20.7 Tuff dacite 0.10 < 124 12 43 59 ATO-TR-238 1.7 70.6 72.3 Fault 0.38 6.0 3980 184 583 12400 ATO-TR-239 0.5 19.6 20.1 Andesite 1.99 < 8420 256 6 97 ATO-TR-240 1.0 19.5 20.5 Andesite 3.16 < 459 332 10 127 1.0 21.5 22.5 Andesite 1.97 < 423 86 6 79 1.2 22.5 23.7 Andesite 1.98 < 1610 218 8 91 ATO-TR-241 1.0 22.5 23.5 Crush zone 0.14 < 685 88 9 258 1.0 41.0 42.0 Andesite 0.33 < 257 81 6 395 ATO-TR-184 1.0 2.0 1.0 Fracture zone 0.11 2.0 218 18 37 222 2.0 4.8 2.8 Rhyolite 0.13 < 321 31 43 112 27.1 28.8 1.7 Rhyolite 0.16 2.0 604 22 27 35 52.8 55.0 2.2 Siltstone 0.12 2.0 152 48 91 80 ATO-TR-187 6.0 6.7 0.7 Silica 0.24 3.0 160 75 107 126 ATO-TR-218 1.0 3.0 2.0 Siltstone 0.13 < 714 194 29 336
Trenching was performed to confirm the results of surface geochemical and grab sampling and the geophysical anomalies that had been surveyed in the area where precious and base metal mineralization found from previous and year 2014.
When excavating trenches, the black-colored topsoil was stripped carefully and piled separately on one side of the trench and 1 m away from the edge of the trench, and the materials below topsoil were dumped on the other side of the trench. Depth of the trench varied depending on the thickness of loose sediments present and it averaged of 2 m. Excavation did not reach the hard rock in some of the trenches where cover sediment was more than 6 meters deep. Width of the trenches varied from 1 to 1.2 m and the walls were slightly sloped to ensure safety. The trenching was aimed at defining alteration zones and soil profiles and evaluating the geochemical and geophysical anomalies that cover large areas. Length of the trenches varied depending on the purpose of each trench, with a minimum of 15 m and a maximum of 276 m. Surveying of trenches were made using differential GPS records.
Documentation of a trench was performed after thorough cleaning of the walls and floor of the trench and it involved mapping at the scale of 1:100 and taking of photos. After that, channel samples and rock chip samples were taken.
After documenting the trench and taking samples, the trench was filled with rock material that was initially taken from the bottom of the trench and covered it with black topsoil material. 100% of the excavated material was put back in the trench and the local authority and local environmental office approved a fact of reclamation. 9.3.3 Geophysical Survey Geophysical data gathered during 2009-2012 was acquired by CGM during its exploration efforts, and included magnetic, gravity, and IP (induced polarization) surveys. Geophysical work included a ground magnetic survey carried out by Monkarotaj, as well as a D-D IP and gravimetric survey completed by Geomaster.
9.3.3.1 Results of Magnetic survey Magnetic surveys were conducted at a grid 25 by 100 m (79. 3 sq.km) south and north of the initial completed grid and also at other prospects in the CGM licensed areas in 2010. Magnetic data were instrumental in identifying major structures, mafic dykes, and ring intrusion-related anomalies. The deposit at ATO is in a
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NI 43-101 TECHNICAL REPORT ATO GOLD PROJECT broad generally low magnetic area that contains linear, low level positive features that coincide with NW and N striking faults (Figure 9.6).
Air-magnetic and spectrometry survey over ~1,000 sq.km (four licenses of CGM). Survey was carried along profiles 100 m apart – 11,021 km in 2011.
Figure 9.6: Magnetic survey of surrounding area of ATO deposit
The result map of the magnetic survey shows that the pipes of the ATO deposit are in an area which generally has weak magnetic intensity with magnetic field values slightly higher to the northwest than to the southeast. In another word, Pipe 4 is characterized by very low magnetic values whereas Pipe 2 has relatively higher magnetic response. A strip of higher magnetic values, which is about 100 meter wide, is observed extending in a northwest direction from the southwest of Pipe 4 to the north of Pipe 1, and this strip spatially coincides with a diorite dyke that is found on the geology map of the area (Figure 9.7).
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Figure 9.7: Detailed magnetic survey of ATO deposit, Au distribution presented.
9.3.3.2 Results of IP Dipole-Dipole (D-D IP) survey D-D IP survey was initially completed across the ATO prospect in the fall of 2009 on recommendations of CGM and then carried on during 2010-2011 using two modifications – 50 and 100 m measurement spacing. D-D IP Sections were placed 100 or 200 m apart – 324.9 km. D-D IP surveys play critical roles in detecting silica and clay alteration (respectively high and low resistivity) and sulfide mineralization (high chargeability), even though a direct 100 percent relationship is not always present. As shown in Figures 9.8-9.9, mineralized pipes at ATO coincide with D-D IP 100-m chargeability and resistivity highs.
Figure 9.8: D-D IP, 100–m chargeability. Dashed blue lines, lines of IP profiles.
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Figure 9.9: D-D IP, 100–m resistivity. Dashed blue lines, lines of IP profiles.
Based on the results in sections, 10 m below the ground surface, several separate areas of very distinct resistivity anomalies were identified and they spatially coincide with a northwest-trending series of small hills that are formed of pipe bodies. Therefore, boundaries of pipes and the locations of trenches are plotted on a map of Level N1, where the anomalies are named after their corresponding pipes (Figure 10). The chargeability map (Figure 11) shows the pipes not so clearly but as some small anomalous areas of medium to weak responses.
Figure 9.10: D-D IP Resistivity map at the ATO deposit. Au distribution presented.
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Figure 9.11: D-D IP Chargeability map at the ATO deposit. Au distribution presented.
Anomalous area of Pipe 1
This anomalous area is about 220 m long and 140 m wide and its size is smaller than the outline of the pipe. Resistivity values range from 100 to 1000 ohm.m. It has been observed that the resistivity values decreases with depth. On the chargeability map, there is a small anomaly that coincides with the northern half of the resistivity anomaly of Pipe 1. With sub-longitudinal strike, this chargeability anomaly is 200 m long and 80 m wide.
The pipe bodies are well distinguished in the dipole-dipole sections of chargeability and resistivity. They are sourced from the large, high response anomalies identified at depth, have moderate values and show columnar and tunnel shapes (Figure 9.12 and 9.13).
Figure 9.12: A cross section for D-D IP Resistivity map at the ATO deposit. Au distribution presented.
Figure 9.13: A cross section for D-D IP Chargeability map at the ATO deposit. Au distribution presented.
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Anomalous area of Pipe 2
Anomalous area measures 280 m long and 130 m wide. Responses of the area are similar to Pipe 1 in terms of values. Chargeability map shows a 350 m long, elongated anomalous area that coincides with the southern part of the resistivity anomaly. They are distinguished clearly in sections (Figure 9.12 and 9.13).
Anomalous area of Pipe 3
This is similar to Pipes 1 and 2 and features an area of high resistivity values. Size of the anomalous area is 300 m by 200 m, relatively larger than the two pipes discussed above. The chargeability map shows three separate anomalies for this anomalous area but they join at depth and form one big anomaly in the dipole- dipole chargeability section. However, on the dipole-dipole resistivity section the anomalous values disappear at depth and this is how the anomalous area of Pipe 2 differs from the other two. Barren rhyolite with pyrite alteration intersected by drill holes serves as the explanation of this anomalous area.
Anomalous area of Pipe 4
No resistivity anomaly is noticed near the ground surface on the resistivity map but only a small area of weak chargeability anomaly is observed in the chargeability map, attracting no interest at this level. However, an anomalous area similar to those of other pipes is manifested in lines of dipole-dipole sections in this part. It was assumed that it may be an underground mineralization after making a comparison of it with the geophysical anomalous areas of the other known mineralization. The small area of weak chargeability anomaly in the dipole-dipole chargeability section has a consistent continuance to depth, is similar to those of Pipes 1 and 2, and is a columnar body sourced from a large anomaly of higher values at depth (Figure 9.14).
Drill hole ATO-111 was drilled to a depth of 500 m in order to check the more extensive and higher geophysical values but it did not yield any mineralization. Therefore, the columnar resistivity and chargeability anomalies of weak to moderate values sourced from the larger anomaly of higher values at more depth serve as a signature for pipe-shaped mineralization.
Figure 9.14: A cross section for D-D IP Chargeability map at the ATO deposit (Pipe 1, 2, and 4). Au distribution presented.
9.3.3.3 Results of Gravity survey A gravity survey (200 by 200 m, 1,704 stations) was completed in 2010 over the ATO prospect and its vicinity. A detailed 50 by 50 m grid was used in the immediate area of Pipes 1, 2 and 3. In 2011, 3,318 stations were completed. The gravity survey of the ATO deposit and its surrounding area shows that the deposit is on the flank of a broad NNE-trending gravity high (Figure 9.15).
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Figure 9.15: Gravity survey of surrounding area of ATO deposit
As seen from the gravity map, the pipes are located within the limit of values of 202.2-203.2 mgal that extend sub-latitudinally. The gravity contours are bent and folded around the ATO deposit and it was interesting that the mineralized pipes are located in these folds (Figure 9.16).
Pipe 1 sits 100 to 150 m wide and is at the center of an area of relatively high gravity response intruding from south to north while Pipe 2 locates at the northern end of a 100 m wide area of relatively high gravity response that slightly pushes from southwest to northeast. Pipe 4 is at the eastern part of an area of relatively high gravity response protruding from south to north.
Figure 9.16: Gravity map at the ATO deposit. Au distribution presented.
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9.3.4 Groundwater Water Exploration Results A comprehensive preliminary hydrogeological exploration program was successfully completed across the
Davkhar basin at ATO prospect area through a 2011 and 2012 work program to support a C2 groundwater reserve estimate under Mongolian classification systems. The exploration program comprised surface geophysics (VES and MRS), core exploration bores completed as observation bores and test bores for aquifer testing.
Exploration works have identified two potential groundwater flow systems extending over an area up to 140km2 that can be characterized as:
Partial coverage of a shallow system extending from surface to between 0 (absent) to 100 mbgl comprising unconsolidated Quaternary and/or weathered Cretaceous sediments. A deeper system extending from 0m to over 250 mbgl comprising semi consolidated Cretaceous sediments.
No continuous low permeability strata has been identified separating the deep and shallow systems across the basin and aquifer testing has demonstrated hydraulic continuity across the systems indicating a leaky or unconfined aquifer system prevails across the exploration area. As such it is assumed the groundwater systems will act as one system through long term supply development. A total saturated sedimentary sequence of at least 250 m has been identified.
Aquifer test interpretation indicates variable transmissivity and aquifer responses from the two test bores with higher transmissivity in the north (>150 m2/d) and lower transmissivity conditions (50 m2/d) in the south. An aquifer storage range of 2.0E-04 – 1.5E-03 has been established from the preliminary aquifer testing. Local barrier boundary conditions are apparent in the south and vertical leakage effects were observed in the north. Individual bore yields of 16 and 18 L/s with pumping bore drawdown limited to 37 and 39 m support the supply potential of the basin. Insufficient data currently exists to firmly characterize permeability distribution which will require assessment as part of future work programs.
Groundwater quality analysis has established various water types across the area which can be generally grouped into ion exchanged waters and areas subject to increased mixing and recharge influences. Groundwater quality is generally brackish across the exploration area; this water should be suitable for industrial purposes but treatment would be required should drinking water use be required.
A numerical groundwater model has been developed to support the groundwater reserve estimate. The model was developed to simulate a conceptual borefield in the southern basin area in proximity to the mine site operating over a 20 year period assuming conservative aquifer parameters. Model predictions indicate a maximum drawdown of 113 m (45% of the initial model saturated aquifer thickness) under assumed worst case reduced hydraulic conductivity and aquifer storage conditions.
A C2 reserve estimate of 417 L/s is presented based on traditional analytical methodology and supported by a preliminary groundwater numerical model with adoption of conservative parameters for each of aquifer area (140 km2), saturated aquifer thickness (150 m) and specific yield (2.5%).
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10 DRILLING
10.1 Drill Programs
The drilling program was completed by CGM during 2010 to 2014. Core drill holes are the principal source of geological and grade data for ATO. Some reverse circulation (RC) drilling and core drilling were completed between 2012 and 2014 through cover to map bedrock geochemical patterns as a method of exploring for blind ATO-style mineralization on the project area. CGM carried out hydrogeology and geotechnical drilling program in 2011 to study related purposes.
End of the 2014, a total of approximately 63,866 m of exploration drilling in 597 holes were completed on the project (Table 10.1). Of this, 54,425 m was core drilling in 370 holes and 9,441 m was completed in 227 RC holes. The drilling has been spread over the ATO mining license as well as south exploration area (Figure 10.1).
Drilling at ATO deposit was conducted along the lines oriented W-NW to E-SE direction and generated 18 cross sections for Pipe 1, 2 and 4 drilling results (see Figure 7.4 presented in Section 7).
Table 10.1: Exploration Drill Hole Summary Table
Exploration Drilling Program Core Holes RC Holes TOTAL
2010 Number 62 62
Length (m) 11,606 11,606
2011 Number 141 141
Length (m) 24,874 24,874 2012 Number 52 90 142 Length (m) 10,444 2,259 12,703 2013 Number 7 137 144 Length (m) 1,539 7,182 8,721
2014 Number 108 108
Length (m) 5,962 5,962 TOTAL Number 370 227 597 Length (m) 54,425 9,441 63,866
10.2 Drill Orientations
The drill holes were mostly drilled at the azimuth of 125 (305) degrees and 60 degrees of inclination through the mineralized pipes. Most of the shallow pattern holes surrounding deposit were drilled vertically. Drilling is normally oriented perpendicular to the strike of the mineralization. Depending on the dipping of the drill hole and the dipping of the mineralization, drill intercept widths are typically greater than true widths.
Average drill hole lengths at the ATO deposit was 186 m. 34% of the drill holes was drilled at the length of more than 200 m where maximum length hole was 716 m.
The drill spacing was on 30 by 30 m and 30 by 60 m grids at the ATO deposit. Drill spacing typically widens toward the margins of the deposits.
Reflex ACT II Rapid Descent tool was used for core orientation.
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Figure 10.1: Location map of drill holes 10.3 Drill Contractors
Diamond and RC drilling on the project were done by Falcon Drilling Mongolia, based out of Canada, which was used BBS-56 rig. Core diameter at the ATO project was HQ-size (63.5 mm nominal core diameter).
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10.4 Core Transport and Storage
At the drill rig, the drillers removed core from the core barrel and place it directly in wooden core boxes. Individual drill runs were identified with small wooden blocks, where the depth (m) and drill hole numbers are recorded. Unsampled core was never left unattended at the rig; boxes were transported to the CGM core logging facility at the ATO camp under a geologist’s supervision. Core was logged on-site before cut for sampling. Remaining core after sampling was transported in sealed boxes by truck to the core shed in Ulaanbaatar.
All core was stored in a secure location in Ulaanbaatar. Core storing workshop was facilitated with the stable shelfs and logging and sample cutting areas. A full time assistant were working for the core workshop to ensure core and samples security and tidiness. 10.5 Geological Logging
Core logging was done in indoor and outdoor for both places in the ATO camp. Upon arrival at the camp core shed, the core was subject to the following procedures:
Quick review. Box labelling check: The core boxes are checked to ensure they are appropriately identified with the drill hole number, meters from–to, and box number written with a permanent marker on the front. Core re-building: Core was usually rotated to fit the ends of the adjoining broken pieces. Core photography. Geotechnical logging, using pre-established codes and logging forms, includes: length of core run, recovered:drilled ratio, rock quality designation (RQD), and maximum length, structural data, and oriented core data. Geological logging: Logging was completed on a paper logging forms. After that, information entered into MS Excel software, which uses standardized templates and validated logging codes that must be filled out prior to log completion. The template includes header information, lithology description and lithology code, graphic log, coded mineralization, and alteration. The geologist marks a single, unbiased cutting line along the entire length of the core for further processing. RC logging involves capture of geological, alteration, and mineralization data on paper logging forms using samples collected in plastic chip trays. 10.6 Core Recoveries
The methodology used for measuring recovery was standard industry practice. In general, core recoveries averaging 97% for the deposit. In localized areas of faulting and/or fracturing, the recoveries decrease; however, this occurs in a very small percentage of the overall mineralized zones. 10.7 Collar and Downhole Surveys
Upon completion of a drill hole, the collar and anchor rods were removed, and a PVC pipe was inserted into the hole. The drill hole collar was marked by a cement block inscribed with the drill hole number (e.g., ATO- 99). Proposed drill hole collars were surveyed by a hand-held GPS unit for preliminary interpretations. After the hole was completed, a differential GPS was used for a final survey pickup. The two collar readings were compared, and if any significant differences were noted the collar was re-surveyed; otherwise the final survey was adopted as the final collar reading.
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CGM used downhole survey instrument, Reflex Instrument AB, to collect the azimuth and inclination at each 50 m in depths of the core drill holes for most of the diamond drill holes. 10.8 Core Sampling
Core samples were taken from diamond drill holes to examine mineralization at depth. Before sampling, core sample was placed in wooden boxes in a proper order, geotechnical measurements were made, and core recovery was estimated. After this, geological documentation and descriptions were recorded and samples were taken with intervals of 1 to 2 m at mineralized zones and 2 to 3 m at unaltered host rocks. The samples were weighed 8-12 kg each.
At the head of the core boxes, notes were put down as to drill hole ID, box number, length of core in the box, and depth intervals in meters. Relevant notes were put down on aluminum plates nailed next to sampled intervals, and the cores in the boxes were then sliced into two by diamond saw. Photo documentation was performed both before and after slicing a sample.
Saw blade was cleaned by working 5 cm deep into a barren rock before slicing a sample. Coolant water was applied in a continuous flow to prevent contamination of the sample during the core slicing process.
After slicing the core, half of every sample was placed in a special plastic bag, which was then tied with cable ties, to be sent to a laboratory. The other half was put back in the box, which was then sealed and sent to Ulaanbaatar for storage.
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11 SAMPLE PREPERATION, ANALYSES AND SECURITY
11.1 Sample Preparation
Sample preparation is a process that requires such delicate and meticulous procedures as implemented in laboratory analyses. Therefore, to avoid loss of volatiles from geochemical samples, the latter was dried in a regular temperature, sieved through 80 mesh sifters, and mixed in a sample splitter to produce 150 to 200 g materials, which then were put in special bags and numbered to be ready for submission to a laboratory. This process of drying, sieving and bagging was done in a sample preparation workshop in Ulaanbaatar, and the remainders of the samples are kept in storage.
Rock samples (grab, chip, channel, and core), they were crushed and prepared in a crushing workshop of an outfit of Actlabs Asia Laboratory (ACTLABS) in the province of Dornod. Before sample preparation, the samples are dried. The actual preparation starts with a jaw crusher Rhino-Terminator, which produces an average of 2 to 3 mm fractions. About 750-1500 g of material from the jaw crusher is then fed into a Lab Tech Essa 2018 type rotary mill, and after that, into bowl and ring pulverizers like Lab tech LM1 and LM2, reducing 90% of the material to 75 μm. After preparation of each sample, all crusher and mills are cleaned by blowing with high-pressure air. Samples each weighing 300 g were sent to an ACTLABS in Ulaanbaatar for assays for precious and base metals. Below is a diagram showing the sample preparation procedure at ACTLABS in Dornod, inclusive of an internal quality control regime. (Figure 11.1)
Figure 11.1: Sample preparation procedure
Sample preparation Quality Control (QC) regimes:
Sieving for remnant samples - 100 g sample was sampled from a crushed material by scoop - Its weight was measured and the symbolized as (A) - It must then be sieved by mesh size of 80 - (+) fractions from sieving was weighed and symbolized as (B) - (-) fractions from sieving was weighed and symbolized as (C) - Percentage of under 80 mesh material was calculated and entered into a spreadsheet and log sheet Sieving for pulverized sample - Scoop about 50 g sample from pulverized material - Its weight was measured and symbolized as (D) - It must then be sieved in mesh size of 150 Page 78 of 209
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- (+) fractions were weighed and symbolized as (E) - (-) fractions were weighed and symbolized as (F) - Percentage of under 150 mesh material was calculated and entered into an spreadsheet and log sheet Control samples of preparatory stage - Duplicate sample was submitted 1 in 50 samples - It must be duplicated from the preceding sample - This will serve to control the preparation process and must be prepared the same way as primary samples Control samples of pulverizing process - A control sample was prepared 1 in 30 samples - It must the duplicated from the preceding sample - This will serve to control the pulverizing process and must be prepared the same way as primary samples 11.2 Laboratory Analyses
All of the core samples, stream sediment samples, soil samples and some of the channel and grab samples were assayed at the ACTLABS in Ulaanbaatar for major elements such as Au, Ag, Cd, Cu, Mo, Pb and Zn. ICP analyses were conducted for a suite of 45 elements in the Stuart Global laboratory. A few of the channel samples and grab samples were assayed in the SGS laboratory (SGS) as a quality control measure.
ACTLABS conducts gold assays within the detection limits of 0.01 to 5 ppm using the FA-ASS method, and performed silver assays within the detection limits of 0.2 to 150 ppm, copper assays within 1 to 10,000 ppm, lead assays within 1 to 8,000 ppm, and zinc assays within 1 to 10,000 ppm, using the AR-AAS method. If there was a grade higher than these limits, gravitation method in ppm was used for gold and silver assays, and AR- ASSAY-AAS method in percentage was used for Cu, Pb and Zn.
SGS performed gold assays within the limits of 0.01 to 1,000 ppm using the FAA method, silver assays within 1 to 1,000 ppm, copper within 2 to 10,000 ppm, lead within 3 to 5,000 ppm and zinc within 2 to 10,000 ppm, using the AAS21R method. When there is a higher grade than these limits, the AAS22S method in percentage is used.
Duration for one submission of laboratory analyses was seven days turnaround during the field exploration and the results were returned periodically. 11.3 Quality Assurance and Quality Control (QAQC)
Quality assurance and quality control of samples are some of the procedures that are mandatory in mineral exploration projects from the discovery of a deposit to a pre-production feasibility study stage. These have many advantages, and most important of these advantages are, firstly, actual existence of a mineral deposit and the assurance to the investors as to quality of work performed by geologists, secondly, controlling of laboratories that perform analyses of samples and materials by the exploration company and geologists, and thirdly, confirmation to the professional regulatory organizations as to the reality of information and data collected by exploration activities.
A quality control procedure was maintained before exploration program. Standards for quality control of sampling and assaying have been well maintained during the exploration work of the company. The following types of control samples have been implemented and these constitute the primary actions of CGM designed to control the works of laboratories.
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2. Duplicate samples taken during sampling (one per 20 to 25 samples) 3. Samples sent to a secondary laboratory for control assays
During the exploration program 2010-2014, a total of 35,821 samples were submitted from drill cores to the laboratories for analyses. 5,180 QAQC samples were analyzed as part of QAQC procedures, and they account approximately 15% of all core samples (Table 11.1). Note that 2012-2014 drilling program conducted about 30% of the drill holes used for resource estimate. The database received for resource estimate in May 2017 was not complete for QAQC result and available data for 2012-2014 was compiled for analysis to addition to the 2010-2011 QAQC analysis by CGM.
Table 11.1: Number of samples for quality assurance and quality control of the ATO project
Core Standard samples Duplicate External laboratory Year samples Au Base Metals samples control samples 2010-2011 24,708 850 606 1,030 2,052 2012-2014 11,113 199 199 237 Нийт 35,821 1,049 805 1,274 2,052
11.3.1 Blank Samples 116 blank samples with no gold grades, which had been prepared in the MEG laboratory of USA, were used in implementation of quality control on the analyses of drill core samples. The results of these samples are compared in Figure 11.2, based on assays from the ACTLABS.
Figure 11.2: Assay results of gold blank samples
Gold grades in blank standard samples are acceptable if they are less than 0.01 g/t. As seen from the graph above, two standard samples gave gold grades higher than that and this is considered as random errors of laboratory. They account for 1.7% of the total blank standard samples. 11.3.2 Standard Samples Standard samples for Au assay that have been used in QAQC were internationally accepted standard samples that were prepared in and imported from the Ore Research and Exploration laboratory of Australia, Geostats Laboratory of Western Australia, Rocklab laboratory of New Zealand, and MEG laboratory of USA, and they totaled 850 samples of 27 kinds in 2010-2011 drilling program. 7 New standards were additionally used for 2012-2014 drilling program at ATO but not analyzed due to a lack of information for both analytical and standard samples (Tables 11.2, 11.3 and 11.4)
For polymetallic elements, QAQC was implemented with 805 standard samples of six different types obtained from the Geostats laboratory. 10 new standard samples were additionally used for 2012-2014 QAQC program but not analyzed because of the same reason mentioned above.
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One standard sample was inserted in every 25 to 30 samples for a quality control measurement. CGM maintained regular reviewing of laboratory results of control samples and standard samples, and compared them with other results and assessed them during the drilling program. If there is a discrepancy between the assayed value and the reference value exceeding ±10% of the latter, an assessment was made either of the laboratory that conducted the assay or of the reference value that was provided by the reference laboratory based on the number of repeats of the discrepancy.
Each time results of assays were received from the laboratory, grades of assayed standard samples were checked against their expected values, and when there were more than two incidences of discrepancies that exceed the maximum allowable amount in assays of gold and other metals in the same submission, then re- assays of all samples of that submission were re-submitted. Such events occurred in assays of samples from drill holes ATO-106, ATO-115, and ATO-156, and all samples from the drill holes were re-assayed.
In addition, statistical processing was performed for every of the metals in the assays of standard samples, and random errors were calculated to account for 0.6-6.6% of the samples. This is thought to be within the acceptable range, and the analytical results were then used in resource estimation process.
The assay result analysis of standard samples by elements are presented in below that were used during the exploration program.
In 2010, 181 standard samples with three different gold grades were obtained from the Ore Research and Exploration Company of Australia (Table 11.2). Their control results returned from the ACTLABS are summarized in Figure 11.3.
Table 11.2: Standard samples from the laboratory of Ore Research and Exploration
№ Standard sample Au, ppm StDEV Type Confidence limits Qty 1 OREAS 5Pb 0.098 0.002 81 2 OREAS 2Pd 0.885 0.029 Oxidized greywacke 0.014 57 3 OREAS 15g 0.527 0.023 Alkaline olivine basalt 0.011 43 Number of Standard samples 181
Figure 11.3: Assay results of standard samples from Ore Research & Exploration
As seen from the control assay results, each of the three grade categories gave two incidences of gold grades outside of allowable limits. However, they only account for 3.3% of total number of standard samples used. Therefore, it was deemed that the assay result of samples from the drill holes that were controlled by these standards may be used in resource estimation.
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Also, additional seventeen standard samples of three different types were obtained from the Geostats Company of Western Australia and used in implementation of quality control on the analyses of drill core samples. (Table 11.3)
Table 11.3: Standard samples from the laboratory of Geostats
Standard № Au ppm StDEV Rock type Confidence limits Qty sample 1 GLG303-1 0.164 16.45 Basalt 3.47 6 2 GLG307-4 0.052 3.21 Granite 1.16 5 3 GLG907-5 0.082 9.17 Oxidized material 2.89 6 Number of Standard samples 17
According to the plot of results of the ACTLABS analyses, these standard samples displayed in Figure 11.4, all grades in standard samples of GLG307-4 and GLG907-5 types are universally close to, or sometimes higher than, maximum random errors while values of GLG303-1 type standard samples show only one such incidence. Therefore, standard samples of GLG307-4 and GLG907-5 types were considered as having higher gold grades than the reference values, and they were no longer used.
Figure 11.4: Assay results of low-gold grade standard samples of Geostats
Quality control on gold assays in 2010 to 2011, 536 control samples of 20 different types were used from the Rocklab Company in New Zealand (Table 11.4).
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Table 11.4: Rocklab Standard samples
Standard № Au ppm StDEV Rock type Confidence limits Qty sample 1 OxA71 0.0849 0.0056 Oxidized 0.0022 41 2 OxC88 0.203 0.01 Oxidized 0.003 45 3 OxD87 0.417 0.013 Oxidized 0.004 35 4 OxE86 0.613 0.021 Oxidized 0.007 32 5 OxF85 0.805 0.025 Oxidized 0.008 25 6 OxG84 0.922 0.033 Oxidized 0.01 33 7 OxG83 1.002 0.027 Oxidized 0.009 31 8 OxH82 1.278 0.029 Oxidized 0.01 33 9 OxI81 1.807 0.033 Oxidized 0.011 26 10 OxJ68 2.342 0.064 Oxidized 0.025 33 11 OxK79 3.532 0.078 Oxidized 0.026 20 12 OxG70 1.007 Oxidized 0.013 5 13 SE-58 0.607 0.019 Sulfide (2.45%S) 0.006 33 14 SF45 0.848 0.028 Sulfide (2.7% S) 0.01 32 15 SG40 0.976 0.022 Sulfide (2.8% S) 0.009 30 16 SH41 1.344 0.041 Sulfide (2.8% S) 0.015 28 17 Si54 1.78 0.034 Sulfide (3.0%S) 0.011 24 18 SN50 8.685 0.18 Sulfide (3.3 % S) 0.062 16 19 Hisilc2 3.474 Siliceous (1.0 % S) 0.034 2 20 HiSilp1 12.05 0.33 Siliceous (2.0% S) 0.13 12 Number of Standard samples 536
Figures 11.5 to 11.8 provide comparisons of reference grades of standard samples against assay results from ACTLABS.
Figure 11.5: Rocklab assay results of standard samples on low-grade oxidized mineralization
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Figure 11.6: Rocklab assay results of standard samples on medium-grade oxidized mineralization
Figure 11.7: Rocklab assay results of standard samples on medium-grade fresh mineralization
Figure 11.8: Rocklab assay results of standard samples on high-grade fresh mineralization
As seen from the above graphs, standards such as SH41, Hisilp, and SN50 each have one random error (2.8%, 1.2% and 1.6%, respectively), and standard OxC88 has three random errors out of 45 samples. Other suites of standard samples have returned grades within acceptable ranges. Only 1.8% of all gold standard samples exceeded the acceptable ranges; therefore, it was concluded that gold assays of primary samples by this laboratory were of good quality.
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As for polymetallic elements, QAQC have been implemented with 606 standard samples of six different types obtained from the Geostats laboratory. (Table 11.5)
Table 11.5: Standard samples of polymetallic elements obtained from Geostats Pty
Standard GBM310-1 GBM399-6 GBM903-7 GBM906-10 GBM301-3 samples
- - - - -
+ + + + +
StDEV StDEV StDEV StDEV StDEV
Metals
Assay ppm Assay ppm Assay ppm Assay ppm Assay ppm Assay
Range of random errors of errors random Range of errors random Range of errors random Range of errors random Range of errors random Range
Ag 19 1.7 0.4 15.5 1.1 0.3 3 0.3 0.1 1.6 0.4 0.1 0.9 0.6 0.1 samples Blank Cu 5825 222 51.4 21373 739 178.8 313 18 3.1 1916 81 20.2 776 51 8.6 Pb 3046 246 60.5 1446 57 15.3 49 5 0.9 1252 82 21.6 11 7 1.4 Zn 9772 540 126.8 2488 164 41 669 36 6.2 595 50 12.6 94 29 5 As 362 34 8.8 175 20 5.4 33 4 0.9 18 3 1.1 304 32 6.1 Ni 37 10 2.4 20 6 1.7 21 4 0.7 32 14 3.7 7910 366 69.9 Co 36 8 2 4 2 0.7 12 3 0.5 27 12 3.4 187 16 3 Number of 149 108 106 140 45 58 Standards
Figure 11.9: Blank standard results for Ag
Figure 11.10: Low-grade standard sample results for Ag
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Figure 11.11: High-grade standard sample results for Ag
Assayed silver grades in blank standard samples are considered acceptable when they are lower than 0.1 g/t, and in this case, only one sample has shown random error (1.7%). (Figure 11.9)
As per other standard samples that contain silver grades, random errors were identified in four incidences. In addition, assayed silver grades in standards GBM906-10 and GBM903-7 are generally close to the high limits of random errors, which is a sign that analyses of ACTLABS tend to yield slightly higher grades of silver (Figure 11.10 and 11.11). Random errors in the standard samples of GBM310-1 type account for 0.6%, in GBM903-7 for 1.8%, and GBM906-10 for 0.7%, and all errors in all types account for 0.8% of the total number of standard samples.
Figure 11.12: Blank standard results for Cu
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Figure 11.14: High-grade standard sample results for Cu
Because copper grades in all blank samples were assayed below 25 g/t, they are considered within the acceptable range (Figure 11.12). Although there is one random error in each of the standards GBM906-1 (0.7 %), GBM310-1 (0.6 %), and GBM399-6 (0.9 %), and the combined random errors for copper account for 0.5% of all of the standards. (Figures 11.13 and 11.14)
Figure 11.15: Blank standard results for Pb
Figure 11.16: Low-grade standard sample results for Pb
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Figure 11.17: High-grade standard sample results for Pb
Assayed lead grades in blank standard samples are considered acceptable when they are lower than 25 g/t, and the graphs above show that all blank samples returned grades within the acceptable range (Figure 11.15). However, there are random errors little higher than the upper limits in three to four samples of each of the standards such as GBM399-6 (4.6 %) and GBM310-1 (2.6%), and one random error slightly out of the acceptable range in a sample of the standard GBM399-6 (Figures 11.16 and 11.17). These random errors account for 1.3% of the total standard samples for lead.
Figure 11.18: Blank standard results for Zn
Figure 11.19: Low-grade standard sample results for Zn
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Figure 11.20: High-grade standard sample results for Zn
Assay grades of zinc lower than 30 g/t are considered acceptable in blank standard, and two samples of this standard returned random errors (3.4% of all blank samples) (Figure 11.18). There was one random error in each of the standards GBM301-3 (2.2%) and GBM399-6 (0.9%) that contain grades of zinc. They account for 0.7% of the total standard samples of zinc (Figures 11.19 and 11.20).
As seen from the graphs above, percentage of random errors of silver and base metals in each of their standards varies from 0.6% to 4.6% and in the total number of their standard samples from 0.5% to 1.3%. Therefore, we believe that the assay results of primary samples returned from ACTLABS can reasonably be used in estimation of resources and reserves of the ATO deposit. 11.3.3 Duplicate Samples During the implementation of the geological and exploration project in 2010-2011, a total of 1030 duplicate samples were taken, accounting for 4.17% of the total number of samples. Duplicate samples were inserted at a frequency of one in 20 to one in 25 primary samples. After a primary sample was cut from drill core, half of the remaining half, or ¼ of the core, was taken as a duplicate. The interval of a duplicate sample was not numbered with the same number as its primary sample to hide their identification; instead, they were given numbers in the order of Dup-1, Dup-2, Dup-3, etc.
Core samples were submitted to the ACTLABS laboratory. A comparison between duplicate and primary core samples is provided in Table 11.6 and Figures 11.21, 11.23, 11.25, 11.27, and 11.29.
Table 11.6: Comparison of primary and duplicate samples
Elements Primary < Duplicate Primary > Duplicate Acceptable range (+50 %) Au 35 3.40% 17 1.65% 978 94.95% Ag 55 5.34% 31 3.01% 914 91.65% Cu 74 7.18% 52 5.05% 904 87.77% Pb 144 14.0% 69 6.70% 817 79.30% Zn 114 11.07% 61 5.92% 855 83.01%
As seen from the table above, duplicate samples tend to have greater elemental grades than do primary samples for each of the elements listed.
Because duplicate samples are taken from the half of the remaining half of a core after the latter has been sampled; therefore, their assay results may have greater deviation from the results of primary samples than do other types of control samples. Because of this we chose the acceptable limits of deviation or variation to be at ±50%.
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Results of primary and duplicate samples and a comparison of their variations are provided in Table 11.7 and in graphs 11.22, 11.24, 11.26, 11.28 and 11.30.
Table 11.7: Comparison of variations of primary and duplicates samples assay results
Variation limits Elements 5% 10% 20% 25% 50% >+50% Au 371 97 181 58 195 128 36.02 % 9.42 % 17.57 % 5.63 % 18.93 % 12.43 % Ag 261 122 217 96 181 153 25.34 % 11.85 % 21.07 % 9.32 % 17.57 % 14.85 % Cu 182 147 235 84 211 171 17.67 % 14.27 % 22.81 % 8.16 % 20.49 % 16.60 % Pb 136 118 221 77 246 232 13.20 % 11.46 % 21.46 % 7.48 % 23.88 % 22.52 % Zn 231 147 204 73 200 175 22.43 % 14.27 % 19.80 % 7.09 % 19.42 % 16.99 %
According to Table 11.6, the highest values belong to gold assay results of the primary and duplicate samples, which supports the conclusion that gold has relatively consistent distribution. However, primary and duplicate samples assayed for lead gave quite fluctuant results.
Figure 11.21: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 20 ppm Au
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Figure 11.22: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 20 ppm Au
As seen from the comparisons between assay results of primary and duplicate samples for gold, they have a correlation of R^2=0.882 (Figure 11.21). Samples other than 52 samples that exceeded the acceptable limit had gold grades less than 0.1 g/t (Figure 11.22). Therefore, 94.95% of the total samples were thought to be within the acceptable range.
Figure 11.23: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 80 ppm Ag
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Figure 11.24: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 80 ppm Ag
As seen from the comparisons between assay results of primary and duplicate samples for silver, there is a correlation of R^2=0.943 (Figure 11.23), which implies the analyses were good. Samples other than 86 samples that exceeded the acceptable range of ±50% had silver grades less than 0.1 g/t, and 91.65% of the total samples were thought to be within the acceptable range, or were low-grade samples (Figure 11.24).
Figure 11.25: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 6,000 ppm Cu
Figure 11.26: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 6,000 ppm Cu
Correlation between copper grades in duplicate and primary samples is R^2=0.86 (Figure 11.25). Samples other than 126 samples that exceeded the acceptable limits had gold grades less than 25 g/t. Therefore, 87.77% of the total samples were thought to be within the acceptable range (Figure 11.26).
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Figure 11.27: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 30,000 ppm Pb
Figure 11.28: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 30,000 ppm Pb
As seen from the comparisons between assay results of primary and duplicate samples for lead, there is a correlation of R^2=0.887 (Figure 11.27). Samples other than 213 samples that exceeded the acceptable range had lead grades less than 30 g/t. Thus, 79.30% of the total samples were thought to be within the acceptable range, or were low-grade samples (Figure 11.28).
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Figure 11.29: Comparison of primary and duplicate sample assay results. Note: a. All samples, b. Assay capped by 70,000 ppm Zn
Figure 11.30: Relation between primary samples and variations. Note: a. All samples, b. Assay capped by 70,000 ppm Zn
Duplicate samples other than 175 samples that exceeded the limits of ±50%, or 83.01% of all samples were thought to be within the acceptable range (Figures 11.29 and 11.30).
As seen from the assay results of duplicate samples shown above, there is a correlation of R^2=0.86-0.943. Also, 79.30-94.95% of the total samples were within the acceptable range, thus the results from exploration drill holes can be used in resource estimation. 11.3.4 QAQC for Laboratories 11.3.4.1 Laboratories Internal QAQC Laboratories that assayed the samples from the ATO deposit implemented their own internal quality control measures by assaying all of the standard samples, gold-blank samples, and other control samples. In particular, they conducted a control assay on remainder of one pulverized sample out of ten. Quality control of assays of all samples have been monitored and graphed. It is considered a warning sign when standard deviation reaches a point two times greater than actual, and control measures are taken when it is three times greater. Re-assays were conducted on selected samples when standards exceeded the warning level, and when the results of re-assays exceeded it again, all samples in the batch in question were internally re- assayed.
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11.3.4.2 External control and validation of laboratory analyses QAQC Samples from holes drilled in 2010-2014 were assayed in the ACTLABS. To conduct control assays on the quality of this laboratory, pulverized materials from the storage were assayed in the SGS. 15 drill holes were selected from drill holes that were located at the core of the orebodies and that showed discrepancies in control and duplicate samples and uncertainties in assay results. Remaining pulverized materials of 2052 samples from these drill holes were assayed in the SGS.
Because remaining pulverized materials are used in conducting the external control measure, the variation of assays are usually higher than standard samples and lower than duplicates due to uneven distribution of valuable elements. Therefore, we used a limit of ±20% in the variation or acceptable range of errors in the samples when comparing the assay results of the said two laboratories.
Based on the said comparison of assay results of the two laboratories, a total of 113 samples, including 25 for gold, 25 for silver, nine for copper, 52 for lead and 24 for zinc, that exceeded the ±20% limit of acceptable range, were selected and re-assayed in the laboratory of ACTLABS (Table 11.8). This accounts for 5.5% of all control samples.
Table 11.8: Changes in correlation of laboratories ACTLABS and SGS
Elements Correlation of ACTLABS and SGS (R^2) First reading Second reading Au 0.990 0.994 Ag 0.988 0.989 Cu 0.835 0.995 Pb 0.943 0.993 Zn 0.986 0.997
After re-assays, correlation of the laboratories has increased slightly, decreasing the random errors of the ACTLABS.
A comparison of assay results of ACTLABS and SGS is provided in Table 11.9 and Figures 11.31 and 11.32.
Table 11.9: Comparison of assay results of ACTLABS and SGS
Elements Laboratories Unit Average grade High Medium Low Background Total Au ACTLABS g/t 12.34 2.28 0.38 0.035 1.12 SGS 12.27 2.23 0.37 0.029 1.10 Ratio % 0.5 2.3 2.4 15.89 2 Ag ACTLABS g/t 314 48 9 1.28 8 SGS 295 44 8 1.07 7 Ratio % 6 8.0 9 16.32 9 Cu ACTLABS g/t 6499 1650 558 83 566 SGS 6787 1597 559 80 568 Ratio % -4 3 -0.1 4 -0.4 Pb ACTLABS g/t 29859 3940 600 83 7007 SGS 29002 3837 570 78 6809 Ratio % 3 3 5 6 3 Zn ACTLABS g/t 43646 3823 618 136 11024 SGS 44468 3876 637 138 11225 Ratio % -2 -1 -3 -1 -2
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As seen from the table above, ratios of background grades for gold (15.89%) and silver (16.32%) are high, which is due to the minimum detection limit of laboratories. Ratios of grades for other elements have errors ranging from -4 to +9. And, grades from ACTLABS are slightly higher than SGS as for gold, silver and lead and some of the copper grades and all of the zinc grades from the laboratory are slightly lower than SGS, as seen in the table above. However, the differences in grades are no much and within the acceptable range. Therefore, assay results of ACTLABS are thought to be of good quality.
Figure 11.31: Results of gold assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS
As seen from the comparison of assay results of laboratories, the correlation is quite high for gold (R^2=.994), and as per variation, most of the samples are within the ±20 range. (Figure 11.31)
Figure 11.32: Results of silver assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS
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According to diagrams in Figure 11.32, there are small discrepancies in three to four samples with high silver grades assayed by the two laboratories but the variation is within the acceptable range of ±20 for these samples.
Figure 11.33: Results of copper assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS
Copper grades were determined with very good quality by both ACTLABS and SGS laboratories, judging from the correlation of R^2=.995 and the fact that most of the samples are within the ±20 range (Figure 11.33).
Figure 11.34: Results of lead assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS
Correlation between assay results from ACTLABS and SGS have been very good for lead, with a correlation of R^2=.993 and variation values within the ±20 range, as seen from the diagrams in Figure 11.34.
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Figure 11.35: Results of zinc assays of ACTLABS and SGS laboratories. Note: a. Comparison of assay results from laboratories, b. Assay results and variation of ACTLABS
As per zinc, the correlation between the results from the two laboratories is R^2=0.997, which is quite close to 1, and this signifies that confidence in these laboratories is high. The diagram of variation shows that most of the samples are within the acceptable range, as shown in Figure 11.35.
According to the diagrams of variations, samples that are out of the ±20 range contain low grades in most cases as for all elements. This is due to the lowest detection limits of the two laboratories.
It has been concluded that the assay results of the samples from drill holes made in the deposit can be used in resource estimate considering that the samples that were used in quality control were able to represent all samples of the exploration work and that the quality validation of samples collected from the deposit was good, judging from the results of assays of standard samples and duplicate samples as well as external control samples. 11.4 Sample Security
Sample security procedure was mentioned in Section 10. 11.5 Considerations and Conclusions
In the Author QP’s opinion, the QAQC procedure and analysis was well performed based on the information above. The Author QP noted several problematic assays in the database. Most importantly cause of the error was not mentioned and the author could not find any documentation about what action was taken.
A statistical analysis of the QAQC data provided by STEPPE was prepared by CGM and it did not reveal any significant analytical issues. The Author QP is of the opinion that the sample preparation, analysis, QAQC and security protocols used for the ATO Project follow generally accepted industry standards and that the data is valid and of sufficient quality to be used for Mineral Resource estimation.
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12 DATA VERIFICATION
12.1 General
The Author QP conducted a review of the geological digital data supplied by STEPPE through GLOGEX for the ATO Project to ensure no material issues could be found and there was no cause to consider that the data was not accurate. The Author QP’s review included site visits undertaken at 23rd of August and 2nd of October 2017, and a desktop analysis.
GLOGEX supplied to the Author QP digital data including previous exploration database. Most geological information was taken from the ATO exploration result 2010-2011 report prepared by CGM and GLOGEX in 2012. This report was submitted to Mongolian government under its guidelines. This report does not meet NI 43-101 and NI 43-101F guidelines. CGM exploration reports for 2010-2014 were also intensely used for geology sections of this technical report.
The Author QP reviewed all QAQC procedures carried out by CGM including a review of logging, sampling and sample preparation procedures; reviewed all technical data including geophysical and geochemical data; carried out an analysis of the analytical QAQC results; and compared data sets with observations made in the field.
The Author QP is satisfied that QAQC procedures carried out by CGM confirm to generally accepted industry standards and that the data used in this report is reliable. The reviewed drilling database formed the underlying data for the independent NI43-101 Statement of Mineral Resources completed by the Author QP. 12.2 Drill Hole Database Validation
Drill hole database used for Mineral Resource estimation was supplied a digital Excel database with collar, survey, lithology, oxidation and assay data. There was an issue pulling out a final database from the received files took additional time.
The final database was cross checked with locations on the generated maps and sections internally prepared exploration reports by CGM in 2014. Assays were cross checked files PDF and Excel files of original assay certificates from the ACTLABS. The Author QP checked all grades and orientation of the drilling against the original assay certificates and cross sections and found no inconsistencies. Original drill hole database was prepared from the geological log sheets and they were used for a random drill hole cross checks conducted against compiled drill hole data. No significant errors were found.
During this review the Author QP noted almost no inconsistencies in the provided data. Only noted that the late mineralization barren dykes received with some significant grades and it was cross checked in loggings and geological researches to decide what action to take during grade modeling filters. Barren diorite dykes intrudes to the pipes in some places and triangulation was modelled and initialized for each metals low background results based on the statistic information of the diorite.
A review of the drill hole digital data was imported to GEMS6.2 (GEMS) software for estimation. Validation process was included a visual validation on screen in sections and drill hole validation module in GEMS to check errors including:
Down hole survey depths did not exceed the hole depth as reported in the collar table; Assay values did not extend beyond the hole depth quoted in the collar table; and Assay and survey information was checked for duplicate records. Interval for all tables missing data.
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No errors were noted. 12.3 Assessment of Database
The database review conducted by the Author QP shows that STEPPE has supplied a digital database that is largely supported by verified certified assay certificates, original interpreted sections, and some report with technical studies.
Based on the data supplied, the Author QP considers that the analytical data has sufficient accuracy to enable a Mineral Resource estimate for the ATO deposit.
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13 MINERAL PROCESSING AND METALLURGICAL TESTING
Metallurgical tests for processing of ore samples from ATO deposit were conducted at the Central Laboratory of Xstrata Support Process in Canada, "ALS Metallurgy-Ammtec" laboratory in Australia and "Boroo Gold" LLC processing plant in Mongolia.
Metallurgical test samples were collected from the exploration drill core and bulk samples for both oxidized and fresh zones within mineralized pipes of Pipe 1, Pipe 2 and Pipe 3. Metallurgical tests for ore samples included a step-by-step leaching test carried out by the roll-up test (bottle-roll test) and granular ore test. Tests were conducted in 4 phases:
1. Bottle-Roll tests of the Xstrata Process Support Center in Canada (from 4 April 2010 to June 30, 2011) 2. Destruction Testing and Analysis - Canadian Xstrata Process Support (September 2011) 3. Column leach tests with low diameter (75 mm) (conducted in crushed ore) - ALS Metallurgy in AMMTEC laboratory in Perth, Australia (completed in November 2011 in April 2012) 4. Additional analysis of column tests with a large-scale diameter (280mm) - in CIL plant of Boroo Gold LLC in Mongolia - September 2012) 13.1 Xstrata Process Support (XPS) Metallurgical Oxide Testing
Xstrata Process Support (XPS) undertook metallurgical testwork on composites from 5 zones from ATO deposit including Upper Oxide Zone, Upper Transition Zone, Lower Transition Zone, Upper Sulphide Zone, and Lower Sulphide Zone composites along with basic mineralogical characterization of the Upper Oxide, Upper Transition and Lower Sulphide Zone composites. Samples were received at XPS in December 2010 and work commenced shortly thereafter. The program used methods and practices developed at the Xstrata Process Support Centre focused on providing quality and preliminary mineralogical and metallurgical data. Phases of work completed are in accordance with the agreed scope of work issued October 13th 2010 and include the following:
• QEMSCAN analysis • Electron Probe Microanalysis • Polished thin section by QEMSCAN • Gravity Recovery Test (GRG) on Upper Oxide (ATO-01). • Leaching tests on Upper Oxide (ATO-01). • Data interpretation and reporting 13.1.1 Mineralogical results Basic mineralogical characterization has been completed using unsized -6 mesh samples of Oxide, Upper Transition and Sulphide Zones. Samples have been analyzed by QEMSCAN (Quantitative Evaluation of Materials by Scanning Electron Microscope) and EPMA (Electron Probe Micro Analysis). This work is preliminary in nature, and more detailed work on representative samples is recommended. A second set of high grade samples (5 in total), selected directly from the uncrushed drill core were prepared as polished thin sections and also measured via QEMSCAN. The reason for including this second set of samples was to define in-situ textures prior to crushing and to search for Au minerals.
QEMSCAN is an automated system that produces mineral maps (colour coded by mineral), through collection of rapidly acquired energy dispersive x-rays (EDX). These maps describe the texture and mineral associations in each of the samples. In addition to the coloured map, the output of the QEMSCAN measurement includes a quantitative measure of modal mineralogy, mineral grain size, and elemental deportment.
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Quantitative compositional analysis was completed using a Cameca SX-100 Electron Microprobe. Electron Probe Microanalysis (EPMA) produces higher electron beam currents and increased beam stability, coupled with higher resolution wavelength dispersive spectrometry (WDS). These features allow for improved detection limits and accuracy of the resulting analysis. Detection limits can be as low as 200 ppm for some elements which provides trace element compositions within the various mineral species in this study. Resulting detailed compositional data is then input back into the QEMSCAN software, in order to refine the final elemental deportment calculations. Compositional data from microprobe analysis can also be used to update the “Species Identification Program” within QEMSCAN.
13.1.1.1 Modal Mineralogy Mineralogical analysis via QEMSCAN was completed on unsized -6 mesh material for three of the five composites. The bulk modal analysis including all sulphide and non sulphide gangue species is summarized in Figure 13.1 and Table 13.1.
Figure 13.1: Modal Mineralogy of Oxide, Upper Transition and Lower Sulphide Zones
Table 13.1: Modal Mineralogy of Oxide, Upper Transition and Lower Sulphide Zones
Oxide Upper Transition Lower Sulphide Chalcopyrite 0.03 0.13 0.14 Sphalerite 0.03 3.84 2.48 Pyrite 0.23 3.2 4.13 Galena 0.14 1.81 1.13 Other Sulphides 0.05 0.02 0.02 Quartz 81.13 67.86 61.37 Chlorite 8.83 15.51 13.35 Muscovite 1.58 1.26 6.08 Biotite 1.82 3.18 6.35 Other Silicate Gangue 0.3 0.13 0.71 Zn Carbonate 0 0.2 0 Pb Carbonate 0 0.02 0 Pb Sulphates/Phosphates 1.78 0.29 0 Fe Sulphates 1.07 0 0 Carbonates 1.4 2.2 3.8 Other 1.59 0.37 0.43
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13.1.1.2 Grain Size QEMSCAN produces a stereologically corrected estimate of average grain size based on the diameter of an equivalent volume spheres.
Table13.2 summarizes the grain sizes obtained in both 6 mesh material and the thin section analysis. Grain sizes for each individual sulphide is included along with the size of all sulphides combined. The size of this last category is largest because there is often a textural association between different sulphides resulting in larger composite grain sizes. Note that the finer grain sizes in the 6 mesh material compared to the polished thin sections is due to the fact that some crushing has occurred and the in-situ textures have not all been preserved.
Table 13.2: Average Grain Size Measurements in Polished Thin Sections and 6 Mesh Material
All Sulphides Chalcopyrite Sphalerite Pyrite Galena ATO-2-1:pts 461 24 283 47 71 ATO-2-2:pts 116 70 191 62 44 ATO-2-3:pts 188 33 439 26 159 ATO-2-4:pts 726 34 622 75 124 ATO-4-3:pts 196 56 277 33 60 Average High Grade Unbroken Ore 338 43 362 48 92 Oxide : - 6 mesh 13 8 10 10 12 Transition : - 6 mesh 36 17 63 20 28 Unoxide : - 6 mesh 24 18 26 21 22
13.1.1.3 Mineral Composition Analysis Compositional analysis using the SX-100 Cameca microprobe was completed for the Lower Sulphide, Upper Transition and Oxide Zones using the same polished sections prepared for QEMSCAN analysis. Analysis result on Oxide Zone is summarized in Tables 13.3. Values that fall below the analytical detection limits are greyed out. Analytical totals for non-sulphides do not include the hydrous component and oxygen is a calculated value.
Table 13.3: QEMSCAN analysis result on Oxide Zone
S Fe Cu Zn As Ag Au Pb Al Total Chalcocite 25.49 0.33 72.25 0.04 0.07 0.12 -0.1 0.1 0.04 98.45 Galena 13.31 0.04 0 -0.01 0 -0.1 -0 86.9 0 100.3 Pb Carbonate -0.01 0 -0.02 -0.04 -0.02 -0.1 -0 80.53 0.01 80.55 Pyrite 51.1 46.33 0.06 0.01 0.28 0.01 -0 0.47 0.01 98.28 Sphalerite 32.51 0.6 0.06 63.94 0.02 0.01 -0.1 0.11 0.01 97.28
Na Mg Al Si K Ca Ti Cr Mn Fe Cu Zn Pb S As O Total
Biotite 0.02 0.81 10.29 14.61 4.73 0.25 0.03 0.01 0 22.38 0.24 0.12 0.67 0.06 0.27 34.13 88.63
Muscovite 0.06 3.24 15.03 22.87 4.87 0.11 0.06 0 0.07 1.84 0.03 0.59 0.01 0.01 -0.01 43.2 92.02
Fe Clay w Zn -0.15 4.27 5.21 6.33 0.01 0.38 0.01 0 0.14 37.14 0.48 1.91 0.53 0.03 -0.01 25.92 82.38
Chlorite 0 13.3 9.08 16.19 0.02 0.69 0.06 0 0.27 8.93 0.17 0.36 0.21 0.02 -0.01 37.69 87.07
Pb Sulph/Phos w Fe -0.06 0.01 1.64 1.2 0.12 0.06 0 0 0 11.05 5.51 1.05 23.99 3.79 0.02 13.3 61.74
Pb Sulph/Phos w As Fe -0.03 -0.1 2.31 1.34 0.3 0.07 0.02 0 -0.01 17.5 1.33 0.64 29.52 2.56 6.27 15.86 77.76
PbCarb -0.03 -0.02 0.02 0.08 -0.01 6.53 0 -0.01 -0.02 0.24 0.06 0.13 60.02 0.02 0.05 6.92 74.06
Siderite wPb -0.31 0.23 2.62 2.27 0.03 0.76 0.01 0 0.01 50.56 0.66 0.44 1.38 0.08 0.4 20.34 79.78
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13.1.2 Metallurgical Testwork Results 13.1.2.1 Feed Preparation and Head Assay Approximately 130 kg of drill core each from Upper Oxide, Upper Trans, Upper Sulphide, Lower Trans Zone and Lower Sulphide Zone composites was received by XPS in December 2010. A fifteen (15 kg) sample each was extracted from Upper Trans (ATO-02), Upper Sulphide (ATO-03) and Lower Sulphide (ATO-05) for SMC testing. A fifteen (15 kg) sample each was extracted from Upper Trans (ATO-02), Upper Sulphide (ATO-03) and Lower Sulphide (ATO-05) for Bond Rod Work Index testing. A ten (10 kg) sample each was extracted from Upper Trans (ATO-02), Upper Sulphide (ATO-03) and Lower Sulphide (ATO-05) for Bond Rod Work Index testing. The remaining drillcore was stage crushed down to 100% passing 10 mesh (1.7mm) and underwent Odds and Evens Blending to produce two homogenized sample lots. These lots were split into representative 2.0 kg charges.
The following table summarizes the feed grades for the Upper Oxide composite: The Upper Oxide (ATO-01) sample has arsenic and antimony at 0.030% and 0.006%, respectively.
Table 13.4: Oxide (ATO-01) Composite
%Cu %Zn %Pb %Fe %S Au(ppm) Ag(ppm) Cd (%) Average 0.09 0.27 1.45 2.61 0.69 2.35 6.37 0.001 St Dev 0 0.02 0.06 0.04 0.04 0.12 0.26 0 RSD (%) 2.39 5.76 4.44 1.43 6.38 5.08 4.06 6.55
13.1.2.2 Minor Element Analysis Table 13.5 shows the minor elements analysis for the composite of ATO-01.
Table 13.5: Minor Elements Analysis for Composites
Element Units ATO-01 Element Units ATO-01 SiO2 % 92.4 Hg µg/g 0.3 AI2O3 % 3.28 As µg/g 270 Fe2O3 % 3.06 B µg/g 2 MgO % 1.26 Be µg/g 0.46 CaO % 0.16 Bi µg/g < 0.09 K g/t 3560 Cd µg/g 6.5 Ti g/t 426 Co µg/g 0.83 Mn g/t 230 Cu µg/g 550 Cr g/t 90 Ni µg/g 11 V g/t < 80 Pb µg/g 9100 Na g/t 174 Sb µg/g 43 P g/t 1750 Se µg/g < 0.7 CI g/t 85 Sn µg/g < 0.5 Fe2O3 % 0.066 Te µg/g < 0.1 SiO2 % 0.63 Zn µg/g 1600
13.1.3 Gravity Recoverable Gold A sample of ATO-01 was submitted to the Knelson Research and Technology Centre to determine its GRG value. The gravity recoverable gold (GRG) test is based on progressive particle size reduction followed by precious metals recovery using a Knelson concentrator at each stage. The progressive size reduction stages allows for precious metals recovery as they are liberated while minimizing over grinding and smearing. The gravity gold recovery for sample ATO-01 was only 4.8% with a tailings grade of 1.92 g/t Au from a head grade of 2.66 g/t Au. The sample was not amenable to gravity concentration and too low to provide a GRG value.
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13.1.4 Bottle Roll Leach Tests Bottle roll cyanidation tests were performed on the Oxide (ATO-01) composite. The ERD for ATO-01 was Average ERD head grade is 0.09% Cu, 0.27% Zn, 1.45% Pb, 2.35 ppm Au, 6.37 ppm Ag.
A design of experiments (DOE) was used. The NaCN concentrations used were 150, 450 and 750 mg/l. The grind sizes (P80) used were 38, 56 and 75 micrometers. A total of 8 tests were performed. Table 13.6 shows the DOE.
Table 13.6: DOE for Leaching Tests on Upper Oxide (ATO-01)
Cyanide Concentration Grind Test No. (mg/L NaCN) (k80, µm) 1 750 75 2 450 56.5 3 750 38 4 150 38 5 150 75 6 150 38 7 450 56.5 8 750 75
The results show that recovery was not sensitive to grind or cyanide concentration. Cyanide consumption increased with increasing concentration. The calculated head grades were very close to the assayed head grade of 2.35 g/t Au, indicating that there is very little free gold. This is also supported by the gravity recoverable gold test.
Cyanidation recoveries did not meet expectations in this phase. XPS recommends further tests to test the impact of several parameters such pre-aeration, longer leaching time and possibly oxygen addition during cyanidation.
Samples of the leach solution were taken at 2, 4, 24 and 48 hours. The Au recovery was computed at each of these times. Figure 13.2 illustrates the Au recovery as a function of leaching time.
Figure 13.2: Au recovery as a function of leaching time
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From Figure 13.2, the leach kinetics are slower at 150 mg/l NaCN (tests 4, 5 and 6). Also, there is some indication that increased recovery can occur at longer than 48 hours of leaching. This is more pronounced at lower NaCN concentration. Table 13.7 shows the results of the DOE at 48 hours of leaching.
Table 13.7: Results of the DOE at 48 hours of leaching
Cyanide Concentration Grind Calculated Head Recovery Cyanide Consumption Test No. (mg/L NaCN) (k80, µm) Grade (g/t Au) (% Au) (kg/t NaCN) 1 750 75 2.37 77.1 3.23 2 450 56.5 2.41 76.1 1.95 3 750 38 2.29 76.2 6.13 4 150 38 2.39 76.2 1.22 5 150 75 2.48 75.1 1.07 6 150 38 2.32 74.1 1.14 7 450 56.5 2.35 77.8 2.16 8 750 75 2.51 76.1 5.19
From the DOE analysis, the concentration of NaCN and grind size tested did not affect Au recovery at 48 hours of leaching. The maximum Au recovery obtained was 77.8% at 450 mg/l NaCN and P80 56 micrometers. The Au recovery for the tests is not high. One possible reason is that the gold particles may be fine grained requiring finer grinding. Mineralogical analysis of the oxide sample would be required to confirm this. Another possible reason is the arsenic and antimony in the feed.
The initial CN concentration had a significant impact on NaCN consumption after 48 hours of leaching. The NaCN consumption increased with increasing initial NaCN concentration. The grind size did not have an impact on NaCN consumption.
The lowest NaCN consumption at 1.1 g/t was obtained at 150 mg/l NaCN and a grind size of P80 75 micrometers. The highest NaCN consumption was 6.1 kg/t. This consumption was obtained at a NaCN concentration of 750 mg/l and a grind size of P80 38 micrometers. 13.2 G&T Metallurgical Oxide Testing Program
The composites (Table 13.8) tested in this program represented the oxide mineralization of the upper portion of the pipes. This preliminary program was designed to study the ore characteristics and metallurgical response of the oxide zone as part of the overall ATO property. The assessment included mineralogy, ore hardness, gold and silver extraction response to grind size, and multi-stage diagnostic gold leach test work in accordance to the scope of work directed by CGM.
Two composites were constructed for this program of study. The Master Composite 1 (MC 1) was assessed by the following parameters:
- Chemical, mineralogical, and diagnostic characteristics of the feed. - Ore hardness by conducting an abrasion index test, Bond rod mill work index test, and Bond ball mill work index test on the feed. - Gold and silver extraction response, including grind versus extraction.
The Column Leach Composite was utilized for the preliminary heap leach amenability test program.
The composites were constructed from half core samples, under the direction of Centerra Gold Inc (CGI). The Master Composite 1 (MC 1) was created from 40 interval samples, whilst the Column Leach Composite was created from 21 interval samples.
The program commenced in November 2011 and the test work was completed in April 2012. Following the test work, the preparation of this technical report commenced.
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The following technical brief summarizes the key technical points of the program. All of the test data generated through the execution of this program can be reviewed in a series of appendices attached to this brief. The appendices are arranged as follows: 13.2.1 Sampling Samples for the ATO Project were received at G&T Metallurgical Services in two shipments. The first shipment, received October 14, 2011, consisted of 40 individual half core samples. These were combined to generate the Master Composite 1. It was then stage crushed to 20mm, homogenized, and rotary split into sub-samples for subsequent comminution and extraction test work. Representative head assay cuts were removed and assayed for elements of interest for this project. A sub-sample was split and sent to ALS Metallurgy – Ammtec for diagnostic leach test work.
A second shipment, received November 1, 2011, consisted of 21 individual half core samples. These were combined to generate the Column Leach Composite.
The composite was staged crushed to 12.5mm, homogenised, packaged, and sent to ALS Metallurgy – Ammtec for preliminary heap leach amenability test work.
Tables 13.8 and 13.9 display the mass and identification of the samples received.
Table 13.8: Master Composite 1
Drill Hole Number Sampling From, m Sampling To, m Weight (kg) Form ATO 93 38.75 39.75 3.4 Half Core ATO 93 37.75 38.75 2.9 Half Core ATO 93 36.75 37.75 2.2 Half Core ATO 93 35.75 36.75 2.8 Half Core ATO 93 34.75 35.75 2.9 Half Core ATO 93 33.75 34.75 3.4 Half Core ATO 93 32.75 33.75 2.3 Half Core ATO 93 31.75 32.75 3.2 Half Core ATO 93 30.75 31.75 3.3 Half Core ATO 92 29.40 30.40 4.1 Half Core ATO 92 28.40 29.40 3.1 Half Core ATO 92 27.40 28.40 3.8 Half Core ATO 92 26.30 27.40 4 Half Core ATO 92 25.20 26.30 3.7 Half Core ATO 92 24.10 25.20 2.5 Half Core ATO 92 23.00 24.10 3 Half Core ATO 92 21.90 23.00 3.5 Half Core ATO 92 20.80 21.90 3.8 Half Core ATO 80 26.00 27.00 3.3 Half Core ATO 80 25.00 26.00 3.3 Half Core ATO 80 24.00 25.00 3.4 Half Core ATO 80 23.00 24.00 2.9 Half Core ATO 80 22.00 23.00 3.6 Half Core ATO 80 21.00 22.00 3.4 Half Core ATO 80 20.00 21.00 3 Half Core ATO 74 55.55 56.80 5 Half Core ATO 74 54.30 55.55 4.5 Half Core ATO 74 53.05 54.30 4.3 Half Core ATO 74 51.80 53.05 4.4 Half Core ATO 74 50.55 51.80 4.2 Half Core ATO 116 9.40 10.40 2.7 Half Core ATO 116 11.40 12.40 2.7 Half Core ATO 116 10.40 11.40 2.7 Half Core ATO 110 24.80 25.80 3.5 Half Core ATO 110 23.80 24.80 4.1 Half Core ATO 110 21.80 22.80 2.7 Half Core ATO 110 20.80 21.80 2.9 Half Core ATO 110 19.80 20.80 1.4 Half Core ATO 110 18.80 19.80 3.4 Half Core Total 132.4
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Table 13.9: Column Leach Composites
Drill Hole Number Sampling From, m Sampling To, m Weight (kg) Form ATO-35 13.00 14.00 3.6 Half Core ATO-35 14.00 15.00 3.4 Half Core ATO-35 15.00 16.00 3.5 Half Core ATO-35 16.00 17.00 2.4 Half Core ATO-35 17.00 18.00 2.7 Half Core ATO-35 18.00 19.00 3.6 Half Core ATO-35 19.00 20.00 1.8 Half Core ATO-35 20.00 21.00 3.1 Half Core ATO-35 21.00 22.00 3.1 Half Core ATO-35 22.00 22.95 3.2 Half Core ATO-35 22.95 23.95 2.9 Half Core ATO-35 23.95 24.95 3.5 Half Core ATO-35 24.95 25.95 3.3 Half Core ATO-89 16.10 17.30 3.5 Half Core ATO-89 17.30 18.50 4 Half Core ATO-89 18.50 19.70 4 Half Core ATO-89 19.70 20.90 3.6 Half Core ATO-89 20.90 22.10 3.8 Half Core ATO-89 22.10 23.30 3.5 Half Core ATO-89 23.30 24.50 2.1 Half Core ATO-89 24.50 25.70 3.8 Half Core Total 68.4
Master Composite 1
The mineral composition data was generated by conducting a Bulk Mineral Analysis with Liberation (BMAL) on the unsized (nominal 75μm K80) composite sample using QEMSCAN. The mineral composition data is summarized in Table 13.10 and Figure 13.2. The following comments relate to the data:
- Sulphide mineral content was approximately 1.5 percent by weight of the feed. - Pyrite represented about 61 percent by weight of the sulphide minerals. - Pyrite, sphalerite, and galena are mainly associated with gangue, whilst the copper sulphides tend to be mainly associated with the pyrite and to a lesser extent sphalerite and gangue. - The liberation data indicates that, at the 78µm K80 sizing, between 20 and 45 percent of pyrite, sphalerite, galena, and the copper sulphides were liberated. - At these liberation levels, flotation of a sulphide concentrate might be viable. However, the presence of talc and barite might cause dilution of the concentrate if not appropriately controlled. - The initial work completed included mineralogical evaluation of the composite sample. The predominant gangue mineral are silica and silicates (77% of sample). Oxide minerals account for 4% of the identified minerals while sulphides account for 1.5% of the sample.
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Table 13.10: Mineral Composition and Elemental Deportment of the Oxide Composite
Mineral Composition (wt. %) Sulphur Deportment (%) Lead Deportment (%)
Sulphide Minerals Mass Gangue Minerals Mass Sulphur Minerals Mass Lead Minerals Mass
Copper Sulphides 0.1 FePb Oxide 1.4 Copper Sulphides 2.2 Galena 42.9 Galena 0.3 FePb Sulphate 1.1 Galena 5.6 Pyromophite (PbPOx) 5.6 Sphalerite 0.1 Pyromophite (PbPOx) 0.1 Fe Pb Sulphate 9.9 Fe Pb Sulphate 37.4 Pyrite 1.0 Quartz/ 77.3 Sphalerite 5.2 FePb Oxide 14.0 Feldspars Arsenopyrite 0.01 Chlorite 11.7 Pyrite 60.6 Iron Oxides 1.4 Arsenopyrite 0.2 Micas 1.4 Barite 16.3 Talc 1.2 Barite 1.1 Smithstonite 0.3 Other Gangue 1.4 Total 1.5 Total 98.5 Total 100 Total 100
Note: 1) Copper Sulphides includes Chalcopyrite, Chalcocite and Covellite. 2) Other Gangue includes Kaolinite (clay), Calcite, Apatite and trace amounts of garnet and other unresolved mineral species.
Figure 13.2: Mineral distribution by class of association Master composite 1, (78μm k80).
Note: Pyrite includes Arsenopyrite.
Diagnostic Leach Test Work – Master Composite 1
The distribution or deportment of gold in various minerals is determined by a series of selective leaches, usually by increasingly stronger oxidative acid leaches. Between each stage, cyanide leaching is used to extract the released gold. In this study a total of ten analysis stages were carried out for the composite at a grind size of 150μm. A summary of the results can be seen in Table 13.11. This test work was conducted at ALS Metallurgy – Ammtec in Perth, Australia.
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Table 13.11: Multi-stage sequential diagnostic gold leach summary
Recovered Distribution Description Au (g/t) (%) Gravity/Free Cyanidable Gold 1.28 79.3 Carbonate Locked Gold Content 0.14 8.7 Iron Oxide Locked Gold Content 0.04 2.2 Arsenical Mineral Locked Gold Content 0.11 7.1 Pyritic Sulphide Mineral Locked Gold Content 0.007 0.5 Silicate (Gangue) Locked Gold Content 0.036 2.2
The diagnostic results indicate that the bulk of the gold content occurred as free (gravity/cyanidable) old, being 79.3 percent of the total gold content. There was a moderate amount of gold associated with iron oxide minerals and arsenical minerals, being 8.7 percent and 7.1 percent, respectively.
The remainder of the gold content was distributed in various categories, but these were generally quite low concentrations, being less than 0.05 g/t.
Ore Hardness – Master Composite 1
Ore hardness was measured using a standard Bond abrasion index, Bond rod mill work index (BRWI), and bond ball mill work index (BBWI) test procedure. A closing sieve aperture of 106µm was utilized for the Bond ball mill work index determination. The results of the tests are summarized in Table 13.12.
Table 13.12: Ore hardness data
Bond BBWI BBWI Abrasion BRWI Composite Closing kWh/ton Index kWh/ton Screen mm Ai MC 1 106 15.6 0.2802 13
The Bond ball mill work index is nearly identical to the Master Composite ball mill work index. Although the material tested is oxidized it is still classified as moderately hard. The high content of silica and silicates that will not readily oxidize provides the hardness. This is also confirmed by the relatively high abrasion index. The sample is quite competent despite its oxidized state. 13.2.2 Ore Characteristics Chemical composition, mineralogical characteristics, diagnostic leach test work, and ore hardness were all measured. These important parameters, that impact on process design, are discussed further in the following subsections.
13.2.2.1 Chemical Content Using standard chemical assaying techniques, the chemical composition of the composite was determined. The results from the analyses are shown in Table 13.13.
Table 13.13: The results from the chemical analyses
Assays – percent or g/t Composite Fe Pb Zn Au Ag S C As Sb PbOx ZnOx Master Composite 1 2.46 0.74 0.31 1.45 27 0.83 0.08 0.019 <0.001 0.29 0.041 Column Leach * 2.02 2.12 0.7 1.61/1.5 10.5 0.3 <0.03 0.012 0.002 - - Note: Au and Ag assays are reported in g/t, all others in percent. * Assays reported by ALS Ammtec.
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13.2.3 Mineral Distributions Mineral distributions from the composite is shown in Figure 13.3.
100
90
80
70
60
50
40
30
Mineral Distribution (wt. %) (wt. Distribution Mineral 20 Multiphase Binary - Gangue Binary - Pyrite 10 Binary - Sphalerite Binary - Galena 0 Copper Sulphide Galena Sphalerite Pyrite Gangue
Figure 13.3: Mineral Distribution by Class of Association
13.2.4 Recovery Response to Grind Size – Master Composite 1 The preliminary cyanidation test program included tests aimed at investigating the effect that grind size had on gold and silver extraction via cyanidation. Heap leaching was subsequently investigated. The following subsections discuss the results of these tests.
A schematic of the test flowsheet, along with the test results, are displayed in Figure 13.4 and Table 13.14. The results of these tests reveal the following points of interest:
- Gold extraction appears to be insensitive to grind size, with recovery being approximately 80 percent for a range of grind sizes; from 78μm K80 to crush size of K100 3,350μm. - Silver extraction increased markedly, from about 50 percent to 69 percent, when the feed size to the cyanidation was decreased from 3,350µm K100 to 299µm K80. Below this sizing, there was no significant improvement in silver extraction. - Sodium cyanide consumption levels are moderate, being less than 2.0 kg/t. Lime consumption decreases as grind size increases, being 4.3 kg/t at 78μm K80 and 1.6 kg/t at 3,350μm K100.
Figure 13.4: Flowsheet - Master composite 1
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Table 13.14: Test results – Master composite 1
Grind Percent Au Extraction at Time Percent Ag Extraction at Time Consumption (kg/t) Test Size Hours Hours μm 2 6 24 48 2 6 24 48 NaCN Lime 1 78 43.6 62.6 71.5 77 39.6 64.3 69.5 71.7 1.9 4.3 2 102 62.7 72.1 75.5 80.1 43.9 57.4 71.2 75.0 1.7 2.5 4 153 65 71.6 76.1 79.4 46.3 59.1 70.2 74.5 1.4 2.0 5 299 57.5 66.5 77.9 82.2 43.2 54.6 64.7 68.9 1.5 1.5 3 3350 53.8 63.2 75.2 80.9 28.1 37.3 47.2 50.4 1.8 1.6 * 3,350μm was K100 crush size.
Figure 13.5: Rate of extraction
13.3 Heap Leach Amenability Test Work – Column Leach Composite
The Column Leach Composite was subjected to a preliminary heap leach amenability test work program. A crush size of 12.5 mm was selected by the client. This test work was conducted at ALS Metallurgy – Ammtec in Perth, Australia. 13.3.1 Natural Percolation Rate Test Work Percolation rate test work was carried out on the Column Leach Composite at the selected crush size of less than 12.5 mm. This test was carried out on the as-crushed ore without, agglomeration with binding agents such as cement or lime. If percolation is deemed satisfactory the next stage of the program would be the column cyanidation leaching of the ore. However, if percolation is poor, then agglomeration testing is necessary prior to committing to column cyanidation test work.
Table 13.15: Natural percolation summary results
Discharge Liquor Slumpage Sample Test Auto L/min L/m2/hr pH (%) MC 1 1 1.09 13,021 7.6 - 2 0.9 10,689 7.6 0.37
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13.3.2 Column Cyanidation Leach Test Work A 50 kg sub-sample of the Column Leach Composite, crush size less than 12.5 mm, was subjected to a 40 day column cyanidation leach. Test summary of results is shown below.
Table 13.16: Column leach test conditions
Leach Wash Crush Percolation Rate* Sample Test Duration Duration Size (mm) (days) (days) L/min L/m2/hr MC 1 JS1059 <12.5 40 7 0.216 413.2 Note: *This is on final day prior to column being dumped.
Table 13.17: Column leach extraction results
Consumption Gold and Silver Extractions Percent at Day Sample Test Element (kg/t) 1 2 5 10 20 30 40 Final* NaCN Lime MC 1 JS1059 Au 30.9 44.3 59.2 68.6 76 80.3 81.9 82.5 4.6 1.6 Ag 3.2 7.7 19.3 33.4 48 57.4 61.5 62.5 Note: * Forty days of leaching plus seven days of water wash.
The results of the column leach cyanidation reveal the following points of interest:
- Column leach cyanidation test work resulted in an excellent gold extraction level of 81.9 percent after 40 days of leaching. The extraction level of silver at 61.5 percent, would be considered good for a column leach cyanidation test. - Gold dissolution kinetics were relatively rapid with the bulk of exposed cyanidable gold being solubilised within ten days from the start of column leaching process. - Sodium cyanide consumption was relatively high, being 4.6 kg/t; this may be due to the open environment to which the column leach is exposed as compared to the closed environment for the earlier bottle roll leach tests. 13.4 Large-Scale Column Tests
Boroo Gold Mine site metallurgical section received 1300 kg samples from ATO oxidized ore. The composites crushed by 3 categories, such as 12.5 mm, 25 mm and 50 mm. Each column diameters were 280mm.
The natural percolation rate of the composite, 200 L/h/m2, indicated by ALS Metallurgy/G&T Metallurgical that agglomeration was not necessary.
All samples classified and Au, Ag grade determined by ACTLABS in UB before loading to column. Leaching solution PН10,0-10,5 stabilized by hydrated lime Ca(OH)2 after the loaded to column 10 % Cyanide solution added to column ore level to 300-500 mg/l after stabilized solution PН10,0-10,5 Solution cycle rate was 12 L/h/m2.
Column test continue 57 days and recoveries reached maximum level on 35th day. Solution and ore ratio was 2.38-2.6:1 end of the test. After column testing, Au and Ag grade in each class determined by lab. Fresh water added to column for 4 days after test stops.
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Figure 13.6: Large scale column test at Boroo Mine
ATO-198 ATO-199 ATO-200 ATO-201 Total weight, kg 369 329 407 937
79 kg left 621 kg left weight to test, kg 369 329 328 316
BLEND
SPLIT
WEIGHT WEIGHT WEIGHT
CRUSH -50mm CRUSH -25mm CRUSH -12.5mm
SIZE DISTRIBUTION
-50+25mm -25+12.5mm -12.5+6mm -25+12.5mm -12.5+6mm -6+3mm -12.5+6mm -6+3mm -3 mm -6+3mm -3 mm -3 mm
SCREEN SAMPLES TO LAB for ASSAY Au and Ag And Bottle roll Figure 13.7: Large scale column test flowsheet
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Figure 13.8: Large scale column testing process
Table 13.18: Large column test results
Recovery (%) Feed grade Ratio Cyanid in solution, Test Bottle roll pH (g/tn) Column test mg/l 24 h -12.5 мм 1.25 36.8 72.65 2.5 370 10.53 -25 мм 1.34 58.1 72.23 2.6 488 10.75 -50 мм 1.12 42.0 58.47 2.38 628 10.86
Table 13.19: Recovery of large column test
Recovery % (Day) Average recovery Test 5 15 25 35 45 57 % -12,5 мм 39.83 66.69 70.61 71.73 72.58 72.65 72.65 -25 мм 54.01 66.78 69.80 71.15 71.91 72.23 72.23 -50 мм 39.44 51.93 55.32 56.71 57.99 58.47 58.47
Table 13.20: Recovery of large column test
Percolation Rate, Time Percolation Rate, Recovery % Time (Day) 2 Recovery % 2 Test (L/h/m ) (Day) (L/h/m ) Solution/Solid ratio 1 : 1 Solution/Solid ratio 2 : 1 -12,5 мм 70.59 23 0.21 72.58 46 0.2 -25 мм 69.80 24 0.2 71.91 45 0.22 -50 мм 55.32 24 0.2 58.23 47 0.2
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Table 13.21: Loaded carbon
Loaded carbon Settled in carbon Test Carbon, gr Au, g/tn Ag, g/tn Au, mg Ag,mg -12,5 мм 17353 13 4.2 226 73 -25 мм 14754 22 3.4 325 50 -50 мм 10579 25 3.4 264 36
Table 13.22: Size by size distribution and recovery by size fraction
Before testing After testing Recovery % (Ore) Au Ag Au Ag Au Ag Class g/t g/t g/t g/t % % Test: -50мм -50мм +25мм 0.6 5.3 0.2 3.2 68.3% 39.6% -25мм +12.5мм 1.1 6.3 0.5 3.9 52.3% 38.1% -12.5мм +6мм 1.5 10.5 0.5 3.5 69.3% 66.7% -6мм +3мм 1.9 11.9 0.7 3.5 62.4% 70.6% -3мм 5.0 15.5 0.9 4.0 83.0% 74.2% Class Test: -25мм -25мм +12.5мм 1.2 7.3 0.3 3.9 73.5% 46.6% -12.5мм +6мм 1.7 14.5 0.3 5.0 80.4% 65.5% -6мм +3мм 1.8 19.0 0.6 6.3 68.0% 66.8% -3мм 3.1 19.1 0.4 5.3 88.3% 72.3% Class Test: -12.5мм -12.5мм +6мм 1.7 14.6 0.3 4.9 80.1% 66.4% -6мм +3мм 2.1 16.4 0.4 5.6 80.8% 65.9% -3мм 5.0 20.4 0.2 5.5 95.6% 73.0%
Table 13.23: Reagent consumption
Lime, NaCN, Test kg /t kg/t -12,5 мм 3.45 1.48 -25 мм 2.71 1.34 -50 мм 2.75 1.10
Figure 13.9: Gold recovery curve (Testing for 12.5mm)
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Figure 13.10: Gold recovery curve (Testing for 25mm)
Figure 13.11: Gold recovery curve (Testing for 50mm) 13.5 Conclusions
A preliminary assessment metallurgical test program aimed to investigate ore characteristics and cyanidation response of gold mineralized composites from the ATO deposit was undertaken. The composites tested in this program were constructed from half core samples.
The gold grade for the Master Composite 1 was 1.45 g/t, whilst the silver grade was 27 g/t.
Mineralogy, conducted on MC 1, indicates the sulphide mineral content was 1.5 percent by weight of the feed, with pyrite representing about 61 percent by weight of the sulphide minerals. The liberation data indicates between 20 and 45 percent of the sulphides were liberated.
Bond ball and rod mill work index tests indicated the MC 1 had values of 15.6 kWh/t and 13.0 kWh/t, respectively. The Bond abrasion index returned a value of 0.2802.
The initial bottle roll cyanidation tests indicated the gold extraction was insensitive to grind size, being approximately 80 percent after 48 hours. However, silver extraction increased markedly, from about 50 percent to 69 percent, when the feed size to the cyanidation was decreased from 3,350µm K100 to 299µm K80. Below this sizing, there was no significant improvement in silver extraction. A preliminary heap leach amenability test work program was undertaken on the Column Leach Composite at a crush size 12.5mm. The natural percolation rate of the composite, 10,689 L/h/m2, indicated that agglomeration was not necessary. The gold extraction after 40 days of column cyanidation leaching was excellent at 81.9 percent, whilst the silver extraction was good at 61.5 percent.
Large-scale column tests result / Boroo gold /: Elements of oxide ore production are gold and silver.
Results of the Large-scale column tests Boroo gold mining recovery have 72.23% for gold and 64.24% for silver metal. The chemical reagent is 1.34 kg of sodium cyanide and 2.71 kg of lime for 1 t mineralized material.
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14 MINERAL RESOURCE ESTIMATES
The Mineral Resource estimation has been independently completed by the Author QP in accordance with the guidelines CIM Definition Standards (2014). Information contained in this Report is based on information provided to the Author QP through GLOGEX by STEPPE and verified by the Author QP. All statistics and analysis, and Mineral Resource estimations were carried out by the Author QP.
Specific software used for the resource estimation were GEMS6.2 (GEMS) for modeling, resource estimation and volumetric reports, SNOWDEN Supervisor8.7 for exploratory data analysis, spatial data analysis and block model (BM) validation, Leapfrog GEO4.0 for grade shell modelling.
The following subsections describe the methods used for and results of the mineral resource estimation for the ATO deposit. 14.1 Databases
All mineral resource estimation work conducted by the Author QP was based on the data received May 2017 from GLOGEX. Data information and supportive reports were received from several backup sources in different days. Database checkup to choose a right information from the backup files took significant time than expected. 14.1.1 Drill Hole Database An MS Excel drill hole database was compiled by CGM in 2014 and included tabulated information for collar, survey, assay, lithology and oxidation.
The initial database contained 265 diamond drill holes including distant holes from the deposit. Thus, 27 distant holes were excluded from the further estimation stages.
The database for resource estimation contains the records from 238 diamond drill holes for a total of 44,284.2 m and detail records are shown in Table 14.1.
Table 14.1: Drill hole database information
Data Information Record count Collar 238 Survey 1,140 Lithology 5,752 Oxidation 706 Assay 32,791
14.1.2 Bulk Density Data A total of 226 bulk density determinations were completed from 2010 and 2011 drilling program. An MS Excel density data was compiled by CGM and contains sample interval, drill hole number, lithology, oxidation level, specific gravity and bulk density measurement, assays and laboratory job numbers. The bulk density determination 2.11 g/cm3 to 3.63 g/cm3 with a mean of 2.58 g/cm3 (Figure 14.1A).
Bulk density data was classified by sample oxidation level for resource estimation. Thus, oxidized samples returned with an average of 2.46 g/cm3, transition level of samples returned with an average of 2.59 g/cm3 and fresh samples returned with an average of 2.64 g/cm3 (Figure 14.1B, 14.1C, 14.1D).
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Bulk density samples were sampled from all rock types with different grades and alteration of the samples had regular spatial distribution. The Author QP considers these samples could represent the deposit in this estimation.
Figure 14.1: Histogram of Bulk Density of ATO. A: All; B: Oxidized; C: Transition; and D: Fresh samples 14.2 Geology and Grade Shell Modelling
14.2.1 Lithology Modelling Lithology modelling of the ATO deposit was not carried out for all types of lithology in this resource estimation. Only Quaternary sediment surface and late diorite solids were created for further block model calculation (Figure 14.2).
To improve future modeling efforts, the Author QP recommends that the lithology type be less and specific for mining needs. Current lithology database has 24 different lithology codes. Some of those lithology types were similar and combined based on the available information ended up with 10 lithology types for further analyses in this report. However, ore body was not significantly controlled by lithology, therefore only late mineralization barren diorite and quaternary sediment surface were modelled.
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Figure 14.2: Diorite Dyke Model and Grade Shells 14.2.2 Fault Modelling The fault model was created by CGM back in 2012. The new drilling result was not made significant changes for fault interpretation and the fault model remained as it was. There were flat lying trust faults and vertical faults along the pipe walls interpreted and shown in Figure 14.2.
Figure 14.3: Fault Model and Grade shells 14.2.3 Grade Shells Grade shells (Figure 14.2 and 14.3) for all three pipes were created based on the 0.1 g/t Au grade threshold as the issuer requested. Those shells were used for future resource modeling and estimation in this report. Grade shell was created by Leapfrog software based on an implicit modeling method and imported to GEMS for resource estimation to constrain mineralizations. There were only 0.1 g/t outer shells created and it assumed blocks with this lower grade constrain is reasonable for heap leaching. 14.3 Block Model
14.3.1 Block Model Properties The cell size or block size of the ATO block model was selected based on mining selectivity considerations for open pit mining. I was assumed the smallest block size that could be selectivity mined referred to the selective mining unit (SMU), was 5 m x 5 m x 2.5 m. The block model was extended to cover all three mineralized pipes of the ATO deposit and it has been pushed down to a depth of 700 m above Sea level.
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The block model was rotated by 55 degrees along clockwise direction adjusting to the exploration cross sections created based on the deposit geology. Table 14.2 presents the properties of the block model and illustrated in Figure 14.4.
Table 14.2: ATO Block model properties
Block Model Properties X (Columns) Y (Rows) Z (Levels) Origin Coordinates (UTM WGS84, Zone49N) 632,070 5,366,450 1,100 Block Extend (m) 160 220 160 Block Size (m) 5 5 2.5 Rotation (degrees) 55
Figure 14.4: Block model dimension in GEMS 14.3.2 Block Model Variables Definition and Tagging The block model contains variables for the metal grade estimates, density, pipe and lithology identifiers, resource classification and rock types.
The block models were tagged by excavation solid wireframe models of diorite and grade shells, wireframe surfaces of topography, quaternary sediments and rock types (oxidation level). The wireframes were assigned a priority to ensure that the blocks were tagged in the correct order in the case of overlaps.
The tagged block models were validated against the drill hole data in cross sections visually as shown in Figure 14.5.
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Figure 14.5: Rock type block model with drill hole oxide information (Cross section 9SE) 14.4 Deposit Statistics
14.4.1 Compositing Original samples were of varying size, which makes their grades not directly comparable. Hence a first operation that needs to be done was to standardize sample size by drill hole sample compositing. 75% of the samples were sampled less than 1.5 m length (Figure 14.6). Thus, the assays were composited into fixed- length to downhole, with the length of 1.5 m. This is a safeguard against the inclusion of an insufficient amount of dilution during compositing. Provisions were taken that incomplete composites would be maintained as is if they were larger than 1 m or merged with an adjoining downhole composite (if they were shorter).
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Figure 14.6: Histogram of Sample Length at ATO deposit
The composites were checked for spatial correlation with the shells and possible zero composite values. Total composite sample after processing was counted 29,653 records.
Raw and composited samples of Au grade were compared by their statistics and histograms shown in Table 14.3 to 14.5.
Compositing process reduced variation of the raw data and it indicates raw sample distribution error become less in in composite samples. Generally all pipes have similar results in composite data. Histogram shapes have no significant changes and it means overall characteristics of raw and composite data are almost the same and it is good to continue further analysis using composite samples in this estimation.
Table 14.3: Raw and composite sample statistics for Pipe 1
Statistics Raw samples Composite samples
Au Pipe 1
Sample 8,698 6,786 Minimum 0.005 0.005 Maximum 71.08 49.69 Mean 1.36 1.24 Standard deviation 2.55 2.14 CV 1.89 1.72
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Table 14.4: Raw and composite sample statistics for Pipe 2
Statistics Raw samples Composite samples
Au Pipe 2
Sample 903 795 Minimum 0.02 0.03 Maximum 46.88 30.88 Mean 0.61 0.58 Standard deviation 2.69 1.77 CV 3.74 3.07
Table 14.5: Raw and composite sample statistics for Pipe 4
Statistics Raw samples Composite samples
Au Pipe 4
Sample 4,476 3,641 Minimum 0.005 0.005 Maximum 166.33 87.1 Mean 1.11 1.02 Standard deviation 4.87 3.65 CV 4.39 3.58
14.4.2 Exploratory Data Analysis The composite sample data for the ATO deposit was imported into Supervisor 8.7 for analyses. Data descriptive statistics, histograms and cumulative probability plots, box and scatter plots were produced for Au, Ag, Pb and Zn grade composites in each three pipes of the deposit and shown below.
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Table 14.6: Descriptive Statistics for Pipe 1
Au (ppm) Ag (ppm) Pb (ppm) Zn (ppm) Composite Samples 6,786 6,786 6,786 6,786 Minimum 0.005 0.1 1 33 Maximum 49.69 1,578 181,307 294,980 Mean 1.24 7.3 7,357 11,213 Standard deviation 2.14 23.9 11,198 17,500 CV 1.72 3.3 2 2
Table 14.7: Descriptive Statistics for Pipe 2
Au (ppm) Ag (ppm) Pb (ppm) Zn (ppm) Composite Samples 795 795 795 795 Minimum 0.03 0.4 105 119 Maximum 30.88 103.3 146,707 184,640 Mean 0.58 4.81 6,357 8,390 Standard deviation 1.77 6.7 11,635 19,133 CV 3.07 1.39 1.83 2.28
Table 14.8: Descriptive Statistics for Pipe 4
Au (ppm) Ag (ppm) Pb (ppm) Zn (ppm) Composite Samples 3,641 3,641 3,641 3,641 Minimum 0.005 0.1 1 23 Maximum 87.1 820.8 122,680 216,080 Mean 1.02 11.7 2,596 4,514 Standard deviation 3.65 36 7,947 13,867 CV 3.58 3.1 3 3
Analysis of the descriptive statistics indicate that the metals within each pipe appear to have a moderate to high variability. A large range, coefficient of variation (CV) and variances were seen in the base metals. This was further supported when the log histograms are analyzed (Figure 14.7-10).
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Figure 14.7: Au (ppm) Histogram for Pipe 1, Pipe 2 and Pipe 4
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Figure 14.8: Ag (ppm) Histogram for Pipe 1, Pipe 2 and Pipe 4
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Figure 14.9: Pb (ppm) Histogram for Pipe 1, Pipe 2 and Pipe 4
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Figure 14.10: Zn (ppm) Histogram for Pipe 1, Pipe 2 and Pipe 4
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The histograms for all metals have a log-normal distribution. Au grade distribution in Pipe 1 indicates there are possibly 2 distinct distributions to interpret compare to Pipe 2 and Pipe 4. Au grades in Pipe 2 and Pipe 4 have similar characteristics.
Generally metals distributed differently in pipes and these will consider for further analysis in this report. Box and Whisker plots for each metals are shown in Figure 14.11-14.
Figure 14.11: Box and Whisker Plot for Au
Figure 14.12: Box and Whisker Plot for Ag
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Figure 14.13: Box and Whisker Plot for Pb
Figure 14.14: Box and Whisker Plot for Zn
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14.4.3 Grade Capping and Evaluation of Extreme Grades Extreme gold, silver, lead and zinc composited grades were evaluated to determine the appropriate level to which high, outlier assays for all metals should be cut back (capping) to restrict their influence. Coefficient of variations in the descriptive statistics to the mean grades suggest that outlier assays are influencing significantly to all metal estimations.
The evaluation of the extreme grades was completed by interpreting probability plots and mean-variance plots. The extreme grades were examined for each mineralized pipes for all metals individually and Au results are shown in Figure 14.15-17.
Figure 14.15: Au Probability Plot and Mean-Variance Plot for Pipe 1
Figure 14.16: Au Probability Plot and Mean-Variance Plot for Pipe 2
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Figure 14.17: Au Probability Plot and Mean-Variance Plot for Pipe 4
Extreme grade examination plots were competed for Ag, Pb and Zn respectively and results for top capping parameters for all pipes are shown in Table 14.9. Average metal loss factor is 4.5%. Metal loss factor for gold in Pipe 4 indicates higher than other pipes but compering data density and accuracy, and higher cut to other pipes, cupping remained as 20 for further estimates.
Table 14.9: Summary statistics for the capping in Pipes
High % Uncut Cut # % Max Grade Loss Pipe Metal Sample Mean Mean Samples Samples (ppm) Capping Metal (ppm) (ppm) Capped Capped (ppm) Factor Pipe 1 Au 6,786 49.86 1.24 12.5 1.21 34 0.50 2.4 Ag 6,786 1,578 7.3 60 6.8 33 0.49 6.8 Pb 6,786 181,307 7,357 70,000 7,234 35 0.52 1.7 Zn 6,786 294,980 11,213 115,000 11,053 34 0.50 1.4 Pipe 2 Au 795 30.88 0.58 11 0.54 5 0.63 6.9 Ag 795 103.3 4.8 40 4.7 4 0.50 2.1 Pb 795 146,707 6,357 70,000 6,136 4 0.50 3.5 Zn 795 184,640 8,390 125,000 8,266 4 0.50 1.5 Pipe 4 Au 3,641 87.08 1.02 20 0.91 18 0.49 10.8 Ag 3,641 820.8 11.7 240 10.9 18 0.49 6.8 Pb 3,641 122,680 2,596 55,000 2,468 19 0.52 4.9 Zn 3,641 216,080 4,514 100,000 4,269 18 0.49 5.4
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14.4.4 Metals Correlation The correlation of the metals within the deposit are typical of epithermal style systems and it was discussed in detail in Chapter 9.
This section will discuss the correlations of Au, Ag, Pb and Zn assays. The results are shown in Table 14.10.
Table 14.10: Resource Metals Correlation Matrix of the ATO Deposit
ATO Deposit - All Au (ppm) Ag (ppm) Pb (ppm) Zn (ppm) Au (ppm) 1 Ag (ppm) 0.22 1 Pb (ppm) 0.41 0.30 1 Zn (ppm) 0.31 0.17 0.79 1
Pipe 1 Au (ppm) Ag (ppm) Pb (ppm) Zn (ppm) Au (ppm) 1 Ag (ppm) 0.18 1 Pb (ppm) 0.48 0.42 1 Zn (ppm) 0.31 0.15 0.73 1
Pipe 2 Au (ppm) Ag (ppm) Pb (ppm) Zn (ppm) Au (ppm) 1 Ag (ppm) 0.63 1 Pb (ppm) 0.48 0.88 1 Zn (ppm) 0.12 0.55 0.62 1
Pipe 4 Au (ppm) Ag (ppm) Pb (ppm) Zn (ppm) Au (ppm) 1 Ag (ppm) 0.17 1 Pb (ppm) 0.39 0.06 1 Zn (ppm) 0.32 0.06 0.92 1
The correlation matrix suggests that Au is often moderately associated with Pb and Zn in all pipes. However, general Ag association to Au in the deposit is low, Ag association with Au in Pipe 2 was strong. Typical strong correlation of base metals of Pb and Zn indicates the deposit type of mineralization.
Mineralization of each pipes has significant differences like Ag of Pipe 2 and Pip 4 correlations to Pb. Explanation of the differences were explained in Section 7 to Section 9 of this report. 14.5 Geospatial Analysis
Due to the style of mineralization found within the deposit, geospatial analysis was completed for composites within the pipes separately. Understanding the grade continuity and determining extend and orientation of the continuity is achieved through interpreting and modeling the experimental variograms. The experimental variograms were calculated for Au, Ag, Pb and Zn within each pipes in Supervisor8.7 Software.
The model parameters fitted to the Au, Ag, Pb and Zn variograms are provided in Table 14.11 and the variograms are described in Figure 14.18 – 14.21.
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Different search ellipsoids were created for each metals in each pipes based on the variogram analysis. The ellipsoid radiuses from pass 1 were established using the ranges from the geostatistical analysis and the ellipsoid radiuses from pass 2 were fixed at values equivalent to 2x the ranges of the first pass (Pass 1) to interpolate blocks that were not interpolated in the first pass. Third pass (Pass 3) was established using ellipsoid radiuses from Pass 2 doubled to interpolate remaining block from Pass 1 and Pass 2 within wireframes of the pipes.
Table 14.11: Variogram Analysis Parameters
Model Semi-Major Rotation angle ATO Pipe Metal Sill Major Axis Minor Axis Type Axis GEMS ZXZ Pipe 1 Au Spherical 0.24 50 45 15 -120,80,-125 Ag Spherical 0.3 50 60 30 -160,-30,-170 Pb Spherical 0.15 45 66 29 30,-50,90 Zn Spherical 0.18 102 91 35 -140,-30,-110 Pipe 2 Au Spherical 0.26 26 33 20 140,-90,-140 Ag Spherical 0.2 48 54 29 -90,-80,-90 Pb Spherical 0.16 70 70 70 120,-90,-90 Zn Spherical 0.18 56 70 36 -120,-60,-90 Pipe 4 Au Spherical 0.26 60 31 25 -90,-80,-90 Ag Spherical 0.23 65 52 40 90,-70,80 Pb Spherical 0.1 100 90 60 -70,-40,90 Zn Spherical 0.1 200 170 54 -70,-20,90
Figure 14.18: Variogram Models for Au in pipes
Figure 14.19: Variogram Models for Ag in pipes
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Figure 14.20: Variogram Models for Pb in pipes
Figure 14.21: Variogram Models for Zn in pipes
14.6 Grade Block Models
The geostatistical results summarized in above sections provided the parameters to interpolate the grade models using 1.5 m composites from the capped grade data in order to produce the best possible grade estimates for the defined resources in the ATO deposit. The interpolation was run on a GEMS point area workspace extracted from the drill hole dataset.
The interpolation profiles were customized to estimate grades separately for each metals within pipes. The Ordinary Kriging (OK) method was selected for the final resource estimation for all pipes.
OK was in 3 passes, with an increased search ellipsoid in each pass to ensure sufficient samples. The minimum number of samples required within the search neighborhood in order for an estimate to be interpolated and a maximum number was specified. The minimum number of samples from any one drill hole was also specified. The interpolation of blocks within a particular mineralized pipes were constrained to use only composites that fell within that pipe. This constraint ensures that sample values are not smeared across the pipe boundaries. The parameters for interpolations by Passes are as follows:
Pass 1 Minimum of six (6) and maximum of twelve (12) composite samples in the search ellipse Octant search, minimum three (3) octants and maximum three (3) composite samples per octant Pass 2 Minimum of four (4) and maximum of twelve (12) composite samples in the search ellipse Octant search, minimum three (3) octants and maximum three (3) composite samples per octant Page 136 of 209
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Pass 3 Minimum of two (2) and maximum of twelve (12) composite samples in the search ellipse
The estimated Au grade is illustrated on a cross section in Figure 14.22 and in 3D view in Figure 14.23.
Figure 14.22: A Cross Section view showing estimated block grades of the ATO Pipes
Figure 14.23: 3D view of estimated Au block grades of the ATO Pipes
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14.7 Block Model Validation
The validity of grade estimates is subject to every decision made and it is essential to validate the final model to ensure that the estimated grades reflect the input information. The ATO block model was validated throughout several processes on all grade estimates. 14.7.1 Visual validation First step was visual comparisons of block grades and composites in cross section and plan view and validation provided a good correlation. Figure 14.24 shows drill hole composites and Au block model on the Section 9SE for visual validation.
Figure 14.24: Cross section for visual validation on Section 9SE 14.7.2 Grade Swath Plots and Local Bias Checks The estimated OK model, ID model (estimated for validation) and grade composites for the Au, Ag, Pb and Zn estimation pipes in the three principal directions charted by the Trend or Swath plots to assist in validating the ordinary kriged estimates. The plots were computed the mean grades from the OK model within 15 to 20 m distance slices in the east-west and north-south directions and within 7.5 m in elevation slices to compare to the mean of the composites. A complete set of trend plots for Au, Ag, Pb and Zn are provided in Figure 14.25. No significant local biases were found in the analysis.
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Figure 14.25: Swath Plots for Au, Ag, Pb and Zn for the deposit: Red-composites, blue-OK Model, green-ID Model
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14.7.3 Global Mean Statistics Global mean statistics tabulated for the composites, OK and ID blocks for each pipes and are summarized in Table 14.12. The composites and OK block statistics compare fairly well for the deposit but there is a small drops in grades between the composites and OK blocks except Ag grade. These drops are caused by declustering of higher grade composites when estimating OK. ID estimate was conducted for validation to OK as a second method and it gives a bit higher grade comparing to OK.
Table 14.12: Global Mean Table for OK and ID Blocks versus composites
Deposit Metal Composites OK Blocks %Difference ID Blocks Sample Mean (g/t) Blocks Mean (g/t) Comp to OK Blocks Mean (g/t) Pipe 1 Au 6,786 1.21 84,964 1.08 0.11 84,964 1.11 Ag 6.8 6.5 0.04 6.6 Pb 7,234 7,370 -0.02 7,530 Zn 11,053 11,576 -0.05 11,874 Pipe 2 Au 795 0.54 18,062 0.54 0 18,062 0.56 Ag 4.7 4.7 0 4.7 Pb 6,136 6,081 0.01 6,225 Zn 8,266 8,012 0.03 8,078 Pipe 4 Au 3,641 0.91 96,812 0.90 0.01 96,812 0.92 Ag 10.9 10.9 0 11.1 Pb 2,468 2,446 0.01 2,462 Zn 4,269 4,196 0.02 4,242 ATO Au 11,222 1.06 199,838 0.95 0.10 199,838 0.97 Deposit Ag 8.0 8.5 -0.06 8.6 Pb 5,610 4,846 0.14 4,934 Zn 8,655 7,652 0.12 7,806
14.8 Gold Equivalent Calculation
In order to assess the value of the total suite of minerals of economic interests in the mineral inventory, formulae have been developed to calculate gold equivalency (AuEq) based on given prices and recoveries from GLOGEX through the issuer.
The base gold equivalent formula incorporates gold, silver, lead and zinc. The assumed metal prices are $1,306/oz for gold, $21.6/oz for silver, $1,844/t for lead and $1,944/t for zinc. Gold and silver are expressed in block grades in the form of grams per ton (g/t). Lead and zinc are expressed in block grades in the form of percentages (%). Metallurgical recovery for gold, silver, lead and zinc are expressed as percentage where silver, lead and zinc percentages are relative to gold recovery. The unit conversion used in the calculation for g/t to oz/t was 31.103477 and pound to gr was 453.359237.
Under potential mining circumstances, the issuer requested gold and silver to use for oxide material AuEq calculation and gold, silver, lead and zinc to use for fresh material AuEq calculation.
The base formula for the oxide block calculation:
퐴푔 ∗ 퐴푔푅푒푣 퐴푢퐸푞(푂푥𝑖푑푒) = 퐴푢 + 퐴푢푅푒푣
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And the base formula for the fresh block calculation:
(퐴푔 ∗ 퐴푔푅푒푣) + (푃푏 ∗ 푃푏푅푒푣) + (푍푛 ∗ 푍푛푅푒푣) 퐴푢퐸푞(퐹푟푒푠ℎ) = 퐴푢 + 퐴푢푅푒푣 Where:
퐴푢푅푒푣 = (1306.6/31.103477 ∗ 푅푒푐퐴푢)
퐴푔푅푒푣 = (21.6/31.103477 ∗ 푅푒푐퐴푔)
푃푏푅푒푣 = (1,844 ∗ 10−6 ∗ 푅푒푐푃푏)
푍푛푅푒푣 = (1,944 ∗ 10−6 ∗ 푅푒푐푍푛)
푅푒푐퐴푢(푂푥𝑖푑푒) = 0.7; 푅푒푐퐴푢(퐹푟푒푠ℎ) = 0.71
푅푒푐퐴푔(푂푥𝑖푑푒) = 0.4; 푅푒푐퐴푔(퐹푟푒푠ℎ) = 0.75
푅푒푐푃푏(퐹푟푒푠ℎ) = 0.6
푅푒푐푃푏(퐹푟푒푠ℎ) = 0.55 Different metallurgical assumptions for the oxide and fresh materials and these assumptions are based on metallurgical testwork and confirmed and supplied by GLOGEX. 14.9 Mineral Resource Classification
14.9.1 Mineral Resource Classification Model Resource classification was done based on a set of rules involving estimation (OK) passes, distances to nearest composites, and search octant constrains used to inform the block estimates.
Measured Resources are those estimated in the first pass. The ranges utilized in the first pass correspond to those of the first spherical component of the variogram used in the Au interpolations.
Indicated Resources are those estimated in the second kriging pass. The ranges utilized in the second pass variogram components.
Note that although the mineralized pipes generally enclose only material considered to be Measured and Indicated, that there are a small number of blocks that were generated in the third interpolation pass, and are considered as Inferred material.
The Mineral Resource classifications in Section 9SE are shown in Figure 14.26.
Figure 14.26: Mineral Resource classification block model, Pipe 1: Red – Measured; Blue – Indicated; Green – Inferred resources.
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14.9.2 Mineral Resource Classification, Category and Definition The resource classification definitions used for this Report are those published by the Canadian Institute of Mining, Metallurgy and Petroleum in their document “CIM Definition Standards for Mineral Resources and Reserves” (2014).
Measured Mineral Resource: that part of a Mineral Resource for which quantity, grade or quality, densities, shape, physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough to confirm both geological and grade continuity.
Indicated Mineral Resource: that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed.
Inferred Mineral Resource: that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes. 14.9.3 Estimation Given the density of the processed data, the search ellipse criteria, and the specific interpolation parameters, the ATO deposit can be classified as Measured Mineral Resources, Indicated Mineral Resources and Inferred Mineral Resources. The estimate is compliant with CIM standards and guidelines for reporting Mineral Resources and Mineral Reserves.
Classified resources were estimated within the three mineralized pipes of Pipe 1, Pipe 2 and Pipe 4 using above 0.1 g/t AuEq cut-off grade. Table 14.13 shows resource for oxide material and Table 14.14 shows resources for fresh material. Volumetrics for the resource estimate have been constrained using the Quaternary surface as the top surface and the mineralized pipe solids as the bottom constrain.
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Table 14.13: Mineral Resource Estimate for Oxide material in Pipe 1, 2 and 4
Contained Metal AuEq (g/t) Tonnage Au Ag Pipe # Classification Au Ag Grade Group (mt) (g/t) (g/t) (koz) (koz) 0.1 - 0.3 0.15 0.17 3.84 0.8 18.2 0.3 - 0.6 0.31 0.33 6.68 3.4 67.1 0.6 - 0.9 0.30 0.59 9.20 5.6 87.9 Measured 0.9 - 1.2 0.32 0.89 8.96 9.2 92.7 1.2 - 1.5 0.34 1.21 8.60 13.1 93.1 1.5 < 1.35 2.26 10.18 97.7 440.8 Total 2.76 1.46 9.01 129.7 799.8 0.1 - 0.3 0.06 0.16 4.04 0.3 8.4 0.3 - 0.6 0.10 0.35 6.46 1.1 19.8 Pipe 1 0.6 - 0.9 0.08 0.56 9.67 1.5 26.2 Indicated 0.9 - 1.2 0.06 0.84 11.28 1.5 20.3 1.2 - 1.5 0.04 1.15 11.69 1.6 16.1 1.5 < 0.13 2.11 11.26 8.9 47.6 Total 0.47 0.98 9.07 14.9 138.4 0.1 - 0.3 0.00 0.19 3.42 0.0 0.1 0.3 - 0.6 0.00 0.24 5.14 0.0 0.1 Inferred 0.6 < Total 0.00 0.21 3.97 0.0 0.3 0.1 - 0.3 0.11 0.19 3.02 0.7 10.8 0.3 - 0.6 0.17 0.34 4.39 1.9 24.2 0.6 - 0.9 0.05 0.63 5.43 1.1 9.5 Measured 0.9 - 1.2 0.03 0.93 5.20 0.8 4.3 1.2 - 1.5 0.01 1.21 5.83 0.3 1.5 1.5 < 0.03 2.32 8.73 2.5 9.4 Total 0.40 0.56 4.59 7.3 59.7 0.1 - 0.3 0.46 0.19 2.89 2.7 42.4 0.3 - 0.6 0.62 0.35 4.15 7.0 83.0 0.6 - 0.9 0.16 0.61 5.36 3.1 27.1 Pipe 2 Indicated 0.9 - 1.2 0.05 0.95 5.53 1.5 8.8 1.2 - 1.5 0.03 1.25 5.55 1.4 6.2 1.5 < 0.07 2.27 7.05 5.0 15.5 Total 1.39 0.46 4.10 20.7 183.0 0.1 - 0.3 0.09 0.19 3.35 0.6 9.8 0.3 - 0.6 0.14 0.38 3.67 1.7 16.7 0.6 - 0.9 0.06 0.62 3.51 1.1 6.4 Inferred 0.9 - 1.2 0.01 0.96 4.46 0.3 1.4 1.2 - 1.5 0.00 1.30 4.07 0.2 0.6 1.5 < 0.04 2.35 3.95 3.2 5.3 Total 0.34 0.64 3.62 7.1 40.1 0.1 - 0.3 0.03 0.19 3.60 0.2 2.9 0.3 - 0.6 0.10 0.33 6.23 1.1 21.0 0.6 - 0.9 0.09 0.61 7.74 1.7 21.3 Measured 0.9 - 1.2 0.04 0.86 9.93 1.2 13.7 1.2 - 1.5 0.03 1.11 14.13 1.2 15.6 1.5 < 0.15 2.24 30.13 10.9 146.9 Total 0.44 1.14 15.49 16.3 221.5 0.1 - 0.3 0.01 0.19 4.65 0.0 1.2 0.3 - 0.6 0.10 0.31 8.73 1.0 26.9 Pipe 4 0.6 - 0.9 0.12 0.57 10.57 2.2 41.2 Indicated 0.9 - 1.2 0.09 0.79 13.58 2.2 37.4 1.2 - 1.5 0.05 1.03 18.46 1.8 32.6 1.5 < 0.22 2.39 31.47 16.9 222.3 Total 0.59 1.28 19.22 24.1 361.5 0.1 - 0.3 0.00 0.23 3.25 0.0 0.1 0.3 - 0.6 0.00 0.26 9.26 0.0 0.5 Inferred 0.6 - 0.9 0.00 0.23 27.40 0.0 0.5 0.9 < Total 0.00 0.24 11.41 0.0 1.1 Measured + Indicated 6.06 1.09 9.05 213.0 1,763.9 TOTAL Inferred 0.35 0.63 3.69 7.1 41.4
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Table 14.14: Mineral Resource Estimate for Fresh material in Pipe 1, 2 and 4
Contained Metal AuEq (g/t) Tonnage Au Ag Pb Zn Pipe # Classification Au Ag Pb Zn Grade (mt) (g/t) (g/t) (%) (%) (koz) (koz) (mlb) (mlb) Group 0.1 - 0.3 0.03 0.14 1.63 0.07 0.20 0.1 1.3 0.0 0.1 0.3 - 0.6 0.75 0.20 2.29 0.17 0.47 4.9 55.4 2.8 7.8 0.6 - 0.9 1.00 0.31 2.99 0.30 0.74 10.1 95.7 6.7 16.3 Measured 0.9 - 1.2 0.99 0.46 3.66 0.45 0.99 14.8 116.3 9.8 21.6 1.2 - 1.5 0.91 0.60 4.59 0.60 1.25 17.4 134.0 12.1 25.0 1.5 < 4.47 1.65 8.10 1.11 1.93 237.1 1,164.1 109.8 190.7 Total 8.14 1.09 5.99 0.79 1.46 284.3 1,566.9 141.3 261.6 0.1 - 0.3 0.01 0.14 1.23 0.08 0.18 0.0 0.4 0.0 0.0 0.3 - 0.6 0.17 0.22 2.22 0.16 0.45 1.2 12.4 0.6 1.7 0.6 - 0.9 0.24 0.32 3.05 0.30 0.73 2.5 23.2 1.6 3.8 Pipe 1 Indicated 0.9 - 1.2 0.28 0.45 3.37 0.49 1.02 4.1 30.2 3.0 6.3 1.2 - 1.5 0.24 0.56 4.23 0.65 1.29 4.4 33.2 3.5 6.9 1.5 < 0.66 1.28 7.27 1.11 2.13 27.4 155.5 16.3 31.2 Total 1.61 0.77 4.93 0.70 1.41 39.6 254.9 25.0 50.0 0.1 - 0.3 0.3 - 0.6 0.01 0.28 3.75 0.10 0.31 0.1 1.6 0.0 0.1 0.6 - 0.9 0.02 0.37 4.39 0.21 0.61 0.3 3.2 0.1 0.3 Inferred 0.9 - 1.2 0.03 0.49 3.30 0.54 0.97 0.5 3.6 0.4 0.7 1.2 - 1.5 0.04 0.55 4.01 0.74 1.22 0.8 5.8 0.7 1.2 1.5 < 0.06 0.57 6.18 1.06 2.59 1.0 11.3 1.3 3.3 Total 0.17 0.50 4.62 0.69 1.48 2.8 25.5 2.6 5.6 0.1 - 0.3 0.00 0.16 1.15 0.09 0.18 0.0 0.0 0.0 0.0 0.3 - 0.6 0.01 0.20 2.90 0.31 0.40 0.1 1.0 0.1 0.1 0.6 - 0.9 0.03 0.24 2.80 0.34 0.88 0.3 2.9 0.2 0.6 Measured 0.9 - 1.2 0.02 0.39 3.45 0.49 1.10 0.3 2.3 0.2 0.5 1.2 - 1.5 0.02 0.49 4.05 0.67 1.50 0.3 2.6 0.3 0.7 1.5 < 0.08 0.73 6.58 1.18 2.94 2.0 17.8 2.2 5.4 Total 0.17 0.53 4.93 0.82 1.98 2.9 26.6 3.0 7.3 0.1 - 0.3 0.00 0.16 0.92 0.10 0.18 0.0 0.0 0.0 0.0 0.3 - 0.6 0.02 0.19 2.32 0.26 0.53 0.1 1.5 0.1 0.2 0.6 - 0.9 0.05 0.25 3.03 0.34 0.83 0.4 4.7 0.4 0.9 Pipe 2 Indicated 0.9 - 1.2 0.03 0.39 3.91 0.49 1.14 0.4 3.9 0.3 0.8 1.2 - 1.5 0.03 0.45 4.90 0.69 1.56 0.4 4.2 0.4 0.9 1.5 < 0.18 0.95 7.21 1.42 2.86 5.6 42.8 5.8 11.6 Total 0.31 0.69 5.71 1.02 2.11 6.9 57.2 7.0 14.5 0.1 - 0.3 0.3 - 0.6 0.00 0.26 2.59 0.24 0.31 0.0 0.2 0.0 0.0 0.6 - 0.9 0.01 0.28 4.07 0.51 0.65 0.1 1.4 0.1 0.2 Inferred 0.9 - 1.2 0.01 0.44 3.48 0.47 1.11 0.1 0.9 0.1 0.2 1.2 - 1.5 0.01 0.62 3.75 0.59 1.26 0.1 0.7 0.1 0.2 1.5 < 0.09 0.81 13.57 2.00 4.72 2.5 41.0 4.1 9.8 Total 0.12 0.72 11.40 1.67 3.88 2.8 44.2 4.4 10.3 0.1 - 0.3 0.30 0.16 2.32 0.04 0.09 1.6 22.7 0.3 0.6 0.3 - 0.6 1.31 0.26 4.65 0.10 0.19 10.8 196.5 2.9 5.6 0.6 - 0.9 0.86 0.44 6.44 0.18 0.33 12.2 177.9 3.4 6.3 Measured 0.9 - 1.2 0.56 0.64 8.51 0.25 0.44 11.5 152.0 3.0 5.3 1.2 - 1.5 0.41 0.87 11.44 0.28 0.47 11.4 149.3 2.5 4.2 1.5 < 1.49 2.24 10.81 0.57 0.94 107.3 517.9 18.8 30.8 Total 4.93 0.98 7.68 0.28 0.49 154.8 1,216.3 30.9 52.8 0.1 - 0.3 0.40 0.16 2.25 0.05 0.10 2.0 29.1 0.5 0.9 0.3 - 0.6 1.80 0.25 4.44 0.12 0.22 14.4 257.3 4.7 8.6 0.6 - 0.9 1.35 0.42 7.45 0.18 0.33 18.4 324.4 5.5 10.0 Pipe 4 Indicated 0.9 - 1.2 1.03 0.62 11.70 0.21 0.39 20.5 385.9 4.8 8.8 1.2 - 1.5 0.77 0.82 16.59 0.25 0.41 20.1 408.2 4.2 6.9 1.5 < 2.16 2.05 15.99 0.45 0.76 142.4 1,112.3 21.7 36.2 Total 7.52 0.90 10.42 0.25 0.43 217.7 2,517.3 41.3 71.3 0.1 - 0.3 0.05 0.17 1.90 0.05 0.08 0.3 2.9 0.0 0.1 0.3 - 0.6 0.31 0.26 6.42 0.08 0.15 2.6 64.7 0.6 1.1 0.6 - 0.9 0.33 0.43 12.15 0.11 0.20 4.6 129.8 0.8 1.5 Inferred 0.9 - 1.2 0.31 0.61 18.51 0.11 0.21 6.1 185.4 0.8 1.4 1.2 - 1.5 0.24 0.74 28.76 0.10 0.20 5.6 220.9 0.5 1.0 1.5 < 0.47 1.40 32.59 0.44 0.89 21.4 496.8 4.6 9.3 Total 1.72 0.73 19.92 0.19 0.38 40.6 1,100.5 7.3 14.3 Measured + Indicated 22.67 0.97 7.74 0.50 0.92 706.2 5,639.1 248.5 457.5 TOTAL Inferred 2.01 0.71 18.11 0.32 0.68 46.1 1,170.1 14.3 30.2
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Notes for Table 14.13-14: 1. The estimates of Mineral Resources has been estimated by the Author QP. 2. Tables are not for a final statement for Mineral Resources. 3. The Mineral Resources include Mineral Reserves. 4. The contained gold estimates have not been adjusted for metallurgical recoveries. 5. All Mineral Resources figures reported in the tables above represent estimates as at August 21, 2017. 6. Totals may differ due to rounding
14.9.4 Derivation of Cut-off Grades Considering the costs, recoveries, dilutions and the location of the Project and the deposit characteristics the QPs and GLOGEX have concluded that the mineralization is potentially economically extractable via typical open cut mining methods, with the LOM (life of mine) mining cut-off grade estimated to be approximately 0.3 g/t AuEq for oxide material and 1.1 g/t AuEq for fresh (unoxidized) material which is based on the expected mining and processing, and cost parameters shown in the Table 14.15.
Table 14.15: Au Equivalent Cut-Off Grade estimation parameters Mineral Resources
Description Rock Type Unit Gold Silver
G&A Oxide $/t 0.8 Fresh $/t 1.2 Mining cost /Ore and waste/ Oxide $/t 2.42 Fresh $/t 5 Processing Cost (heap leaching) Oxide $/t 5 Processing Cost (milling) Fresh $/t 25.8 Recovery (heap leaching) Oxide % 70 40 Recovery (milling) Fresh % 71 75 Sales Price $/oz 1306.6 21.6 Refining charge $/oz 6 0.6 Royalty % 2.5 5
Gold equivalent (AuEq) Cut-off grades for oxide and fresh ore were calculated for the ATO using following equation:
(푀𝑖푛𝑖푛푔 + 푃푟표푐푒푠푠𝑖푛푔 + 퐺&퐴)[$/푡] 퐶푢푡 푂푓푓 (퐴푢퐸푞 푔/푡) = (푀푒푡푎푙 푃푟𝑖푐푒 − 푅푒푓𝑖푛𝑖푛푔 퐶표푠푡 − 푀푒푡푎푙 푃푟𝑖푐푒 ∗ 푅표푦푎푙푡푦)[$/푔] × 푅푒푐표푣푒푟푦
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14.9.5 Mineral Resource Statement The ATO deposit Mineral Resource Summary for oxide and fresh materials are presented in Table 14.16 and 14.17.
Table 14.16: ATO Mineral Resource Summary for Oxide material
Contained Metal Tonnage Au Ag Classification Au Ag (mt) (g/t) (g/t) (koz) (koz) Pipe 1 (0.3 g/t AuEq Cut-off) Oxide Measured 2.61 1.53 9.30 129 782 Indicated 0.41 1.11 9.87 15 130 Measured + Indicated 3.02 1.48 9.37 144 912 Inferred 0.00 0.24 5.14 0 0 Pipe 2 (0.3 g/t AuEq Cut-off) Oxide Measured 0.29 0.70 5.19 7 49 Indicated 0.93 0.60 4.69 18 141 Measured + Indicated 1.23 0.62 4.81 25 190 Inferred 0.25 0.80 3.72 7 30 Pipe 4 (0.3 g/t AuEq Cut-off) Oxide Measured 0.42 1.20 16.20 16 219 Indicated 0.58 1.30 19.42 24 360 Measured + Indicated 1.00 1.25 18.06 40 579 Inferred 0.00 0.25 14.11 0 1 ATO Deposit (0.3 g/t AuEq Cut-off) - TOTAL Oxide Measured 3.33 1.42 9.81 152 1,049 Indicated 1.92 0.92 10.22 57 631 Measured + Indicated 5.25 1.23 9.96 208 1,680 Inferred 0.26 0.79 3.81 7 31
Table 14.17: ATO Mineral Resource Summary for Fresh material
Contained Metal Tonnage Au Ag Pb Zn Classification Au Ag Pb Zn (mt) (g/t) (g/t) (%) (%) (koz) (koz) (mlb) (mlb) Pipe 1 (1.1 g/t AuEq Cut-off) Fresh Measured 5.71 1.41 7.30 1.00 1.77 260 1,340 126 224 Indicated 1.01 1.03 6.17 0.94 1.82 33 200 21 41 Measured + Indicated 6.72 1.36 7.13 0.99 1.78 293 1,541 147 264 Inferred 0.12 0.56 4.95 0.87 1.84 2 19 2 5 Pipe 2 (1.1 g/t AuEq Cut-off) Fresh Measured 0.11 0.67 5.96 1.05 2.58 2 21 3 6 Indicated 0.22 0.86 6.79 1.29 2.64 6 48 6 13 Measured + Indicated 0.33 0.80 6.52 1.21 2.62 9 69 9 19 Inferred 0.10 0.79 12.69 1.87 4.42 3 42 4 10 Pipe 4 (1.1 g/t AuEq Cut-off) Fresh Measured 2.06 1.85 10.86 0.49 0.81 122 718 22 37 Indicated 3.24 1.63 15.88 0.38 0.64 169 1,652 27 46 Measured + Indicated 5.29 1.71 13.93 0.42 0.70 292 2,370 50 82 Inferred 0.81 1.12 30.09 0.30 0.60 29 785 5 11 ATO Deposit (1.1 g/t AuEq Cut-off) - TOTAL Fresh Measured 7.88 1.52 8.21 0.87 1.53 385 2,080 150 266 Indicated 4.47 1.46 13.24 0.55 1.01 209 1,900 55 99 Measured + Indicated 12.35 1.50 10.03 0.75 1.34 594 3,980 205 366 Inferred 1.04 1.02 25.44 0.52 1.12 34 847 12 26
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Notes: 1. Estimate of Mineral Resources has been estimated by the Author QP. 2. ATO Mineral Resources are as at August 21, 2017. 3. Mineral Resources include Mineral Reserves. 4. Contained gold estimates have not been adjusted for metallurgical recoveries. 5. Mineral Resources are estimated using a 0.3 g/t AuEq cut-off grade for oxide material and a 1.1 g/t AuEq.cut-off grade for fresh material. 6. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability for heap leaching. 7. A conversion factor of 31.103477 grams per ounce 453.59237 grams per pound are used in the resource estimates. 8. AuEq has been calculated using assumed metal prices ($1,306.6/oz for gold, $21.6/oz for silver, $1,844/t for lead and $1,944/t for zinc). Oxide ore calculation: AuEq(g/t) = Au(g/t) + (Ag(g/t) x 21.6 x 0.0321507 x 0.4) / (1,306.6 x 0.0321507 x 0.7) Fresh ore calculation: AuEq(g/t) = Au(g/t) + ((Ag(g/t) x 21.6 x 0.0321507 x 0.75) + (Pb% x 1,844 x 10-6 x 0.6) + (Zn% x 1,944 x 10-6 x 0.55)) / (1,306.6 x 0.0321507 x 0.71) 9. Totals may not match due to rounding.
14.10 Factors That Could Affect the Mineral Resource Estimate
Areas of uncertainty that could materially affect the Mineral Resource estimates include the following in general:
Commodity pricing. Interpretations of fault geometries. Effect of piping system as a control on mineralization. Lithological and structure interpretations on a local scale, including dyke modelling. Pit slope angles. Metal recovery assumptions. Dilution considerations. Estimates of operating costs used to support reasonable prospects assessment. Changes to drill spacing and number of drill hole composites used to support classification categories. The QPs are not aware that there is a known environmental, permitting, legal, title, taxation, social- economic, marketing or political issue to materially affect the resource estimation. The environmental permit granted to CGM in 2011 is currently under update by STEPPE and will be finalized before mine commencement permit by the Fourth Quarter of 2018.
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15 MINERAL RESERVE ESTIMATES
The Mineral Reserves for the ATO project include the Pipe 1, Pipe 2, and Pipe 4 oxide zones only. The Mineral Reserves were estimated as of August 21, 2017 by the Author QP and reviewed by Mining QP using economic parameters and ultimate pit design by GLOGEX.
Mining, geotechnical, and hydrological factors have been considered in the estimation of the Mineral Reserves, including the application of dilution and ore recovery factors, where appropriate. The QPs notes that other modifying factors (such as metallurgical, environmental, social, political, legal, marketing, and economic factors) have also been considered to the required standard, and that they each demonstrate the viability of Mineral Reserves in their own regard.
The Mineral Reserve definition is done by first identifying ultimate pit limits using economic and geometrical parameters with pit optimization techniques. The resulting optimized pit shells were then used for guidance in pit design to allow access for equipment and personnel. Several phases of mining were defined to enhance the economics of the project, and GLOGEX used the phased pit designs to define the production schedule to be used for cash-flow analysis for the feasibility study. 15.1 Pit optimization
Pit optimization was done using Gemcom Whittle software (version 4.6) by GLOGEX to define pit limits with input for economic and slope parameters. Optimization used only Measured and Indicated Resources for oxide ore processing. All material under 0.3 g/t AuEq ore was considered as waste.
Varying gold prices were used to evaluate the sensitivity of the deposit to the price of gold as well as to develop a strategy for optimizing project cash flow. To achieve cash flow optimization, mining phases or push backs were developed using the guidance of Lerchs-Grossman pit shells at lower gold prices. 15.1.1 Economic Parameters Economic parameters for pit optimizations were used based on the current information. The recoveries are based on the metallurgical test works. Economic parameters are provided in Table 15.1.
Table 15.1: Pit optimization Economic Parameters
Description Unit Value Reference Mining cost /Ore and waste/ $/t Mined 2.42 Processing Cost $/t processed 5.00 G&A $/t processed 0.80 Throughput Rate for oxide ore processing t/year 1,200,000 Days per Year of Processing days/year 330 Gold royalty % of gold price 2.5 Silver royalty % of silver price 5.0 Refining charge of Gold $/oz 6.00 Refining charge of Silver $/oz 0.60 Sales Price, Gold $/oz 1306.6 Sales Price, Silver $/oz 21.6 Initial capital cost $ million 19.4 Gold recovery % 70 Silver recovery % 40 Ore loss % 2.0 Ore dilution % 3.0 CUTOFF GRADE (oxide ore zone AuEq) g/t of oxide ore 0.3
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Mining and processing costs are reflective of the final costs determined for this study. General and administrative (“G&A”) costs are assumed to be fixed at $ 0.8 per ton processed oxide ore.
A base price of $ 1,306.6 per ounce of gold was used for cutoff grade calculations and Whittle project evaluations. Other gold prices were also used for sensitivity analysis. 15.1.2 Cut-off grades Gold equivalent (AuEq) Cut-off grades were calculated for the deposit and Mineral Reserves were estimated and reported using this cut-off. The AuEq Cut-off was calculated as follows:
(푀𝑖푛𝑖푛푔 + 푃푟표푐푒푠푠𝑖푛푔 + 퐺&퐴)[$/푡] 퐶푢푡 푂푓푓 (퐴푢퐸푞 푔/푡) = (퐴푢 [$/푔] × 퐴푢 푅푒푐표푣푒푟푦) + (퐴푔[$/푔] × 퐴푢 푅푒푐표푣푒푟푦)
Where:
퐴푢[$/푔] = (퐴푢 푃푟𝑖푐푒 − 퐴푢 푅푒푓𝑖푛𝑖푛푔 퐶표푠푡 − 퐴푢 푃푟𝑖푐푒 ∗ 퐴푢 푅표푦푎푙푡푦)
퐴푔[$/푔] = (퐴푔 푃푟𝑖푐푒 − 퐴푔 푅푒푓𝑖푛𝑖푛푔 퐶표푠푡 − 퐴푔 푃푟𝑖푐푒 ∗ 퐴푔 푅표푦푎푙푡푦)
Table 15.1 shows the input parameters used in the Cut-off grade calculations for the Mineral Reserve estimates. In this report 0.3 g Au/t cutoff for pit definition, scheduling, and Proven Reserve and Probable Reserve statement. 15.1.3 Slope Parameters Pit slope recommendations were provided by geotechnical assessment of ATO slopes study report in 2011 by Victor Vdovin, the Corporate Geotechnical Engineer, Centerra Gold Inc., Technical Development Group.
GLOGEX used these recommendations to develop parameters for pit design and pit optimization. The recommendations allow for up to a 50 degree inter-ramp angle. 15.1.4 Pit Optimization Result Whittle pit optimizations were run using the economic and slope parameters described in previous sections. Pit optimizations were completed using prices of $950 to $1,650 per ounce Au in increments of $65 per ounce in order to analyze the deposit’s sensitivity to gold prices. Results for $65 per ounce increments from $910 to $1690 per ounce of gold are shown in Table 15.2. A graph of the Whittle results is shown in Figure 15.1.
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Table 15.2: Whittle Pit Optimization Result (AuEq cut-off 0.3 g/t)
Open pit Open pit Mine Mine cash flow cash flow ton Waste life life Final Price factor best worst input best years years pit $K discounted $K discounted best ton best worst 1 0.60 116,914 116,914 4,027.4 1,238.7 3.4 3.4 2 0.65 117,206 116,941 4,126.6 1,317.7 3.4 3.4 3 0.70 117,426 116,803 4,236.3 1,394.0 3.5 3.5 4 0.75 117,526 116,617 4,317.6 1,439.9 3.6 3.6 5 0.80 117,593 116,320 4,422.2 1,520.2 3.7 3.7 6 0.85 117,599 116,051 4,489.2 1,563.6 3.7 3.7 7 0.90 117,584 115,775 4,559.4 1,624.3 3.8 3.8 8 0.95 117,551 115,564 4,604.9 1,663.6 3.8 3.8 9 1.00 117,522 115,425 4,631.3 1,692.9 3.9 3.9 10 1.05 117,448 115,180 4,680.7 1,774.8 3.9 3.9 11 1.10 117,396 115,030 4,708.9 1,831.7 3.9 3.9 12 1.15 117,369 114,962 4,721.2 1,855.0 3.9 3.9 13 1.20 117,350 114,915 4,728.5 1,872.5 3.9 3.9 14 1.25 117,324 114,861 4,737.2 1,893.7 3.9 3.9 15 1.30 117,307 114,827 4,742.2 1,906.2 4.0 4.0 16 1.35 117,294 114,802 4,746.1 1,914.5 4.0 4.0 17 1.40 117,288 114,790 4,747.6 1,917.7 4.0 4.0
Figure 15.1: Graph of Whittle Results (AuEq cut-off 0.3 g/t)
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15.1.5 Ultimate Pit Limit Selection The ultimate pit limit was determined using cash-flow analysis. Whittle software “Pit by Pit” analysis was used based on a 10 discount rate. The resulting $1,300 pit optimization (or pit version-7) was visually compared with the pit designs, and though there are some small portions of designs outside of pit shells, where ramps exist, the pit designs are reasonable with respect to the pit optimization.
Figure 15.2: Ultimate Whittle pit shell and pipes 15.2 Pit design
Detailed pit design was completed, including an ultimate pit. The ultimate pit was designed to allow mining of economic resources identified by Whittle pit optimization while providing safe access for people and equipment. Internal pits or phases within the ultimate pits and were designed to enhance the project by providing higher-value material to the leach pad earlier in the mine life. 15.2.1 Bench Height Pit designs were created to use 2.5 meters benches for mining. This corresponds to the resource model block heights, and GLOGEX believes this to be reasonable with respect to dilution and equipment anticipated to be used in mining. 15.2.2 Pit design Slopes Slope parameters were provided in a report by geotechnical assessment of ATO slopes study report in 2011 by Victor Vdovin, Corporate Geotechnical Engineer, Centerra Gold Inc., Technical Development Group.
Slope recommendations were given as a single set of parameters for all walls in the pit as follows:
Up to 50 degree inter-ramp slope angles; Up to 70 degree bench face angles; Up to 5.0 meter high benches; and Minimum 2.5 meter catch benches.
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15.2.3 Haulage Roads Ramps were designed to have a maximum centerline gradient of 10%. In areas where the ramps may curve along the outside of the pit, the inside gradient may be up to 11% or 12% for short distances. Ramp width was determined as a function of the largest truck width to be used in mine planning. Design criteria accounts for 3.5 times the width of the truck for running room in areas using two-way traffic. An additional width is added to the ramp for a single safety berm at least half of a tire height inside of the pit. For roads designed outside of the pit, and additional safety berm is accounted for in the road widths. Haul road width – 12m. 15.2.4 Ultimate Pit As discussed in previous sections, the ultimate pit limit uses pit shell number 7 from the Whittle runs. The ultimate pit is an integrated design using interior pit phases and it separates 3 small mines. All phases are mined by contract mining. Figure 15.3 and 15.4 show the ultimate pit design.
Figure 15.3: Ultimate pit design (Surpac mine design)
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Figure 15.4: Ultimate pit design 15.3 Dilution and Loss
The Author QP’s grade model with block sizes of 5 m by 5 m by 2.5 m was used to estimate resources. The model was estimated based on this block size, and this model was used to define the ultimate pit limit and to estimate Proven Reserves and Probable Reserves. While the pit design is based on max 2.5 m bench heights, in areas of high/waste grade variability, the benches could be mined at the meter block heights as required. Thus, GLOGEX believes that the block size is reasonable with respect to a selective mining unit and it is not more than 3% and 2%. GLOGEX further believes that this represents an appropriate amount of dilution for statement of reserves for the ATO deposit. 15.4 Mineral Reserves Estimate
Mineral Reserves for the project were developed by applying relevant modifying factors in order to define the economically extractable portions of the resource. The QPs estimated and reviewed the Mineral reserves based on the Cut-off grades by GLOGEX study and the Mineral Reserves meet NI 43-101 standards. The NI 43-101 standards rely on the 2014 CIM Definition Standards on Mineral Resources and Mineral Reserves adopted by the CIM council. CIM standards define modifying factors as:
Modifying Factors: Modifying Factors are considerations used to convert Mineral Resources to Mineral Reserves. These include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.
CIM standards define Proven and Probable Mineral Reserves as:
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Probable Mineral Reserve: A Probable Mineral Reserve is the economically mineable part of an Indicated, and in some circumstances, a Measured Mineral Resource. The confidence in the Modifying Factors applying to a Probable Mineral Reserve is lower than that applying to a Proven Mineral Reserve.
Proven Mineral Reserve: A Proven Mineral Reserve is the economically mineable part of a Measured Mineral Resource. A Proven Mineral Reserve implies a high degree of confidence in the Modifying Factors. 15.5 Mineral Reserves Statement
Table 15.3 reports the Proven and Probable Reserves within oxide ore based ultimate pit designs discussed above. The reserves are shown to be economically viable based on the cash-flow model provided by GLOGEX. The Mining QP has reviewed the cash-flow model and believes that it is reasonable for the statement of the Proven and Probable Reserves for oxide pit reserve. Note that there are some fresh (or transition) material above cut-off grade of 0.3 g/t AuEq inside of the ultimate oxide pits reported as oxide pit reserve in Table 15.3.
Table 15.3: ATO Deposit Mineral Reserves for Oxide Pits
Contained Metal Tonnage Au Ag Classification Au Ag (mt) (g/t) (g/t) (koz) (koz) Pipe 1 (0.3 g/t AuEq Cut-off) Proven 2.74 1.54 9.29 136 818 Probable 0.41 1.13 9.93 15 132 Total Reserve - Pipe 1 3.15 1.49 9.38 151 950 Pipe 2 (0.3 g/t AuEq Cut-off) Proven 0.24 0.74 5.13 6 40 Probable 0.82 0.60 4.70 16 124 Total Reserve - Pipe 2 1.06 0.63 4.80 22 164 Pipe 4 (0.3 g/t AuEq Cut-off) Proven 0.39 1.25 17.22 16 219 Probable 0.57 1.33 19.98 25 367 Total Reserve - Pipe 4 0.97 1.30 18.85 40 586 ATO Deposit (0.3 g/t AuEq Cut-off) Proven 3.37 1.45 9.92 157 1,076 Probable 1.80 0.96 10.74 55 623 TOTAL RESERVE - ATO 5.18 1.28 10.21 213 1,700
Table 15.4 represents the Proven and Probable Reserves including mining dilution and ore loss. The total ore tonnage before dilution and ore loss is estimated at 5.18 Mt at an average grade of 1.28 g/t Au for 213 thousand oz. The dilution tonnage 0.15 Mt represents 3% of the ore tonnage before dilution.
Table 15.4: Mining dilution and Ore loss Reconciliation for ATO Deposit Mineral Reserves
Contained Metal Tonnage Au Ag Classification Au Ag (mt) (g/t) (g/t) (koz) (koz) Ore before ore loss and dilution 5.18 1.28 10.21 213 1,700 Less: Ore loss - 2% 0.10 1.28 10.21 4 34 Ore before mining dilution 5.07 1.28 10.21 209 1,666 Add: Mining dilution - 3% 0.15 0.18 3.22 1 16 TOTAL RESERVE - ATO 5.23 1.24 9.91 210 1,681
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The final ATO Deposit Mineral Reserves represents in the Table 15.5. The total ore tonnage is estimated at 5.23 Mt at an average grade of 1.25 g/t for 147 thousand oz of recovered Au and at an average grade of 10.00 g/t for 673 thouand oz of recovered Ag.
Table 15.5: ATO Deposit Mineral Reserves Statement as at August 21, 2017
Contained Metal Recovered Metal Tonnage Au Ag Classification Au Ag Au Ag (mt) (g/t) (g/t) (koz) (koz) (koz) (koz) ATO Deposit (0.3 g/t AuEq Cut-off) 70% 40% Proven 3.41 1.41 9.72 155 1,065 108 426 Probable 1.82 0.93 10.52 55 616 38 246 TOTAL RESERVE - ATO 5.23 1.25 10.00 210 1,681 147 673 Notes: 1. Estimate of Mineral Resources has been estimated by the QPs. 2. ATO Mineral Reserves are as at August 21, 2017. 3. Mineral Reserves are included in Mineral Resources. 4. Mineral Reserves are constrained within an optimized pit shell based on a gold price of $1,300 per ounce. 5. Mining dilution factor is 3% and Ore loss factor is 2%. 6. Contained gold estimates have been adjusted for metallurgical recoveries. 7. The open pit Mineral Reserves are estimated using a cut-off grade of 0.3 g/t AuEq for oxide material. 8. A conversion factor of 31.103477 grams per ounce 453.59237 grams per pound are used in the reserve and resource estimates. 9. AuEq has been calculated using assumed metal prices ($1,306.6/oz for gold, $21.6/oz). Oxide ore calculation: AuEq(g/t) = Au(g/t) + (Ag(g/t) x 21.6 x 0.0321507 x 0.4) / (1,306.6 x 0.0321507 x 0.7) 10. Totals may not match due to rounding.
The QPs are not aware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that could materially affect the Mineral Reserve estimate. 15.5.1 In-pit Inferred Resources Inferred Resources were not used in the economic analysis. Note that CIM standards define inferred resources as:
An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity.
Table 15.4 shows the Inferred Resources inside of the pit designs for each pipe.
Table 15.6: In-Pit Inferred Resources
Contained Metal Classification Tonnage Au Ag Au Ag Inferred Reserve (mt) (g/t) (g/t) (koz) (koz) Pipe 1 Pipe 2 0.21 0.78 3.71 5 25 Pipe 4 Total Inferred 0.21 0.78 3.71 5 25
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16 MINING METHODS
16.1 Material Types
Material was broken into ore and waste categories for the purpose of scheduling. Waste consists of material that is not included into Proven and Probable Reserves. Ore is categorized by grade, classification, and metallurgical type. 16.1.1 Waste Material Waste material is to be placed into the waste dump immediately. Initial material will be built up around the crusher to provide a stockpiling area. Excess waste will expand the dump to the west to an elevation of approximately 1040 meters to provide for additional stockpiling areas. 16.1.2 Ore Material types Measured and Indicated material above cutoff grade and inside of the pit was classified by both grade and metallurgical domain. The material types were mineralized-grade, low-grade, medium-grade, and high-grade using minimum cut-off grades, 0.50, and 1.0 g/t AuEq grade breaks. The minimum cut-off grades were based on cut-off grades determined for oxide ore type: 0.3 g/t AuEq.
All ore material is considered to be economic and is either feed directly to the crusher or in stockpiles near the crusher. 16.2 Mining Method
The ATO project has been planned as an open-pit truck and shovel operation. The truck and shovel method provides reasonable cost benefits and selectivity for this type of deposit. Only open-pit mining methods are considered for mining at ATO. 16.3 Mine-Waste Facilities
The waste dump is intended to be built as a flat surface extending through the valley to the north. The dump height is not expected to exceed 10 m between the dumping crest to the underlying topography. Waste material would be dumped against a berm on the dump face, and dozers would be used to maintain the dumping face.
The waste dump has a total capacity of more than 20 million tons of the required 2,846 thousand tons of waste, though there is suitable space to expand the dump to the north or upward. 16.4 Mine Production Schedule
Proven and Probable Reserves were used to schedule mine production, and Inferred Resources inside of the pit were considered as waste. The final production schedule uses trucks and shovels as required to produce the ore to be fed into the process plant and maintain stripping requirements for each case.
Table 16.1 shows the mine-production schedule.
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Table 16.1: Mine-production schedule
Total rock Oxide ore tonnage, Waste Years tonnage ROM tonnage Au g/t Ag g/t Au, K Oz Ag, K Oz 2018-IY 544,172 302,204 241,968 1.16 8.80 11.3 86 2019 1,868,323 1,212,138 656,185 1.29 8.66 50.1 337 2020 1,909,434 1,215,459 693,975 1.29 10.45 50.3 408 2021 1,795,449 1,212,138 583,311 1.32 10.25 51.5 400 2022 1,849,953 1,212,138 637,815 1.14 11.26 44.6 439 2023 105,578 73,180 32,399 0.72 4.94 1.7 12 8,072,910 5,227,257 2,845,653 1.25 10.00 209.5 1,681
1,400,000 1,212,138 1,215,459 1,212,138 1,212,138 1,200,000
1,000,000
800,000 693,975 656,185 637,815 583,311 600,000
400,000 302,204 241,968 200,000 73,180 32,399 - 1 2 3 4 5 6
Oxide ore tonnage, ROM Waste tonnage
Figure 16.1: Mine production plan
500 439 450 408 400 400 337 350 300 250 200 150 86 100 50.1 50.3 51.5 44.6 50 11.3 12 1.7 - 2018-IY 2019 2020 2021 2022 2023
Ag, K Oz Au, K Oz
Figure 16.2: Gold and Silver in ROM ore
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16.5 Mining Equipment
The following Section contains equipment selection and fleet requirements in order to carry out the mine plan for the open pit. The mine will be operated with a Contractor fleet which is presented in Table 16.2.
GLOGEX uses Sherpa surface mine software for equipment optimization and fleet selection.
Table 16.2: Mining Equipment Fleet
Fleet size Equipment’s and production segment Equipment Specifications 2018- IY 2019 2020 2021 2022 2023-I Hydraulic Shovel-Ore 2.3 cubic meter 1 1 1 1 1 1 Rear-Dump Truck-Ore 32 metric ton 2 2 2 2 2 2 Front-End Loader-Waste 1.5 cubic meter 1 1 1 1 1 1 Rear-Dump Truck-Waste 32 metric ton 1 1 1 1 1 1 Rotary Drill-Ore 17.15 centimeter 1 1 1 1 1 1 Rotary Drill-Waste 17.15 centimeter 1 1 1 1 1 1 Bulldozers 60 kilowatt 2 2 2 2 2 2 Graders 115 kilowatt 1 1 1 1 1 1 Water Tankers 9500 liter 1 1 1 1 1 1 Tire Service Trucks 1800 gvw 2 2 2 2 2 2 Bulk Trucks/Powder Buggies 275 kilogram/minute 1 1 1 1 1 1 Lighting Plants 7.8 kilowatt 4 4 4 4 4 4 Pick-up Trucks 680 kilogram 3 3 3 3 3 3
Total 21 21 21 21 21 21
16.6 Mining Operations
The mine production schedule is based on a 7 day per week schedule, with two 12 hour shifts per day. There are four crews planned to cover the rotating schedule. Each 12 hour shift contains a half-hour down for blasting and miscellaneous delays, a half-hour for shift start up and shutdown and an hour for lunch breaks for a total of 10 effective working hours.
Each year contains unscheduled time for nine holidays and two non-productive weather shifts, equivalent to 330 days of mine operation per year. Because the mine contract is based on production targets, GLOGEX has not included any equipment availabilities nor utilization rates for any mine equipment in this report 16.6.1 Drilling and Blasting Production drilling is covered in the mine contract, while a separate blasting sub-contract under the contract miner is in place at ATO. STEPPE is responsible for diesel fuel as it is applied to ammonium nitrate and fuel oil (ANFO) blasting agents. 16.6.2 Loading and Hauling The main loading units at ATO are 2.3 m3 hydraulic shovel and 1.5 m3 front-end loader. 32 metric ton haul trucks are the main hauling units; the loaders will require 4 to 5 passes to load the trucks. GLOGEX completed a haulage study to ensure that the mining fleet was sufficient to meet production targets in the mine schedule. Ancillary equipment appears reasonable for an operation of this size and scale.
16.6.2.1 Hauling road Ramps were designed to have a maximum centerline gradient of 10%. In areas where the ramps may curve along the outside of the pit, the inside gradient may be up to 11% for short distances. Ramp width was
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Contract mining has been assumed, using 32 ton capacity trucks. The operating width of the trucks was assumed to be 4.5 m. Ramps are designed to allow 3.5 times the operating width of the trucks along with room for sufficient safety berms. The ramp width used for design is 15 m. 16.6.3 Mine Personnel Mine personnel estimates include both operating and mine-staff personnel for contractor. Operating personnel are estimated as the number of people required to operate trucks, loading equipment, and support equipment to achieve the production schedule. Mine staff is based on the people required for supervision and support of mine production. The estimated number of contract mine operating personnel required is shown in Table 16.3 while the mine-staff requirements are shown in Table 16.4.
Salaries for mine-staffs positions were estimated based on information received from Khishig Arvin, Mongolian biggest contract mining company. Salaries include an allowance for benefits at a rate of 20% of the base salary for each position.
Table 16.3: Contractor’s mine operating personnel Requirements
Workforce year-1 year-2 year-3 year-4 year-5 Driller 1 1 1 1 1 Blaster 2 2 2 2 2 Excavator Operator 3 3 3 3 3 Truck Driver 5 5 5 5 5 Heavy Equipment Operator 5 5 5 5 5 Utility Operator 2 2 2 2 2 Mechanic 3 3 3 3 3 Electrician 1 1 1 1 1 Maintenance workers 4 4 4 4 4 Laborer 2 2 2 2 2 Total work force 28 28 28 28 28 Workers per shift 14 14 14 14 14
Table 16.4: Contractor’s mine-staff requirements and salary
salary per 2018 IY 2019 2020 2021 2022 2023-I Number of month Job tittle total workers of each staff, Dollars/Year USD Mine manager 1 2,500 7,500 30,000 30,000 30,000 30,000 2,500 130,000 Foremen 2 1,500 9,000 36,000 36,000 36,000 36,000 3,000 156,000 Engineer 1 1,500 4,500 18,000 18,000 18,000 18,000 1,500 78,000 Environmental 1 1,500 4,500 18,000 18,000 18,000 18,000 1,500 78,000 Technician 3 1,000 9,000 36,000 36,000 36,000 36,000 3,000 156,000 Purchasing 1 800 2,400 9,600 9,600 9,600 9,600 800 41,600 Secretary 1 700 2,100 8,400 8,400 8,400 8,400 700 36,400 Total 10 39,000 156,000 156,000 156,000 156,000 13,000 676,000 Social insurance-13% 5,070 20,280 20,280 20,280 20,280 1,690 87,880 Total salary and Social insurance 44,070 176,280 176,280 176,280 176,280 14,690 763,880
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17 RECOVERY METHODS
This section describes the recovery methods used for the ATO Project for the crushing, heap leach and process facilities. Flowsheet development, operating parameters and design criteria were based on results from metallurgical test work presented in Section 13. The gold recovery process was designed on the basis of leaching 1.2 Mt of ore per year with an average gold grade of 1.25 g/t, average silver grade of 10.0 g/t at an overall gold recovery of 70%, silver recovery 40%.
The three-stage crushing plant will operate at a nominal 5860 t/d throughput, 275 days per year. The process plant, located near to and down-gradient from the heap leach facility HLF to minimize the pumping and pipeline requirements for pregnant and barren solutions, will operate 365 days per year. The pregnant solution will flow to the plant at a nominal rate of 200 m3/h and a design flowrate of 250 m3/h. The plant is designed to process 5 ton of carbon per day using an adsorption, desorption and refining (ADR) process to extract gold from the pregnant solution to produce the gold doré. The gold ore processing facilities will include the following unit operations:
Crushing and Ore Handling
Primary crusher: a vibrating grizzly screen and jaw crusher in open circuit producing a final product P80 of approximately 190 mm; Secondary and tertiary crusher: a vibrating screen and cone crushers operating in reverse closed circuit producing a final product P80 of 25 mm, and; Heap placement: crushed ore stacked to a 3,000 ton - capacity stockpile, reclaimed by a radial stacker. Heap Leach Pad
Crushed ore stacking Ore leaching; and Barren and pregnant solution delivery and recovery piping systems. ADR Plant
Carbon-in-Column (CIC) Adsorption: adsorption of solution gold onto carbon particles. Desorption: acid wash of carbon to remove inorganic foulants, elution of carbon to produce a gold- silver rich solution, and thermal regeneration of carbon to remove organic foulants, carbon stripping to recover gold into solution. Gold refining: gold electrowinning (sludge production), filtration, drying, mercury retorting, and smelting to produce gold dore. 17.1 Process Design Criteria
The process design criteria and mass balance detail the annual ore production, major flows, and plant availability. Several considerations to mitigate the Tsagaan Ovoo climate, especially with respect to the severe winter conditions, have been included in the general design criteria:
Pregnant solution will be heated as required to maintain a minimum discharge temperature at the heap leach pad of 60C; The drip emitter lines will be buried in the winter to provide a degree of insulation; Barren and pregnant solution–feed pipelines will be buried a minimum depth of 2.5 m; Dedicated standby generators will be installed for backup power supply to the barren solution pumps; The crushing and process plant buildings will be pre-engineered steel structures with insulated steel roofs and walls. Page 160 of 209
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Table 17.1: Design Criteria
CHARACTERISTICS UNITS VALUE Capacity Mtpa 1.2 Ore moisture % 2.5 Ore specific gravity t/m3 2.6 Bond Ball Mill Work Index kwh/t 15.6 Abrasion Index 0.2802
Ore grade Au g/t 1.13 Ag g/t 9.25
Plant recoveries Leach gold recovery % 70 Leach silver recovery % 40
Operating Schedule
Crushing and Stacking Operation Operating Days per Year days 275 Operating Hours per Day h 24 Availability % 75 Operating Hours per Year h 4,950 Feed rate t/h 242.42
Plant Operation Operating Days per Year days 365 Operating Hours per Day h 24 Availability % 90 Operating Hours per Year h 7,884 Gold Production oz/a 30,516.2 Silver Production oz/a 142,743.2
Heap leach Heap leach pad capacity t 5,565.691 Crush size (80% passing) mm 25 Lift Height m 8
Number of Lifts 3
Angle of Repose 37 Stacked ore density t/m3 1.6 Solution Application Rate L/h/m2 12 Leach Time days 60 Liner system 1.5 mm LLDPE on compacted soil base Leak collection and recovery Sump at low point with pipe Leach solution application rate m3/h 200-250
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17.2 Production Rate and Products
A gold and silver production schedule has been prepared based on the mining plan and head grades as discussed in mining section as well as the expected recoveries determined from the metallurgical data review. These are shown in table below.
Table 17.2: Production plan
Gold in ROM Silver in ROM ore, Silver Gold produced, Silver produced, Period Gold recovery ore, oz oz recovery oz oz 2019 49,834 337,583 70% 40% 34,883.56 135,033 2020 49,834 401,884 70% 40% 34,884 160,754 2021 49,834 401,884 70% 40% 34,884 160,754 2022 49,834 401,884 70% 40% 34,884 160,754 2023-I 10,193 138,059 70% 40% 7,135 55,224 Total 209,527 1,681,295 70% 40% 146,669 672,518
17.3 Process Description
A block flow diagram and a schematic flowsheet for the plant and process flows are shown in Figure 17.1 and Figure 17.2.
Figure 17.1: Diagram for the process flows
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Figure 17.2: Schematic flowsheet for the process flows 17.3.1 Crushing A single crushing circuit will be utilized for preparation of the heap leach feed comprised of a conventional closed-circuit jaw / cone / cone crusher flow sheet.
Run-of-mine ore (ROM) will be trucked from the open pits and dumped directly into a primary feed hopper. Primary crusher feed will be drawn from the feed hopper by a feeder discharging onto a vibrating grizzly screen. The grizzly screen oversize will feed the primary jaw crusher. The grizzly undersize and jaw crusher product are to be transported to the secondary screen by a secondary screen feed conveyor, which is equipped with a metal detector and magnet.
The crushing plant will operate 275 days per year. If the crushing plant is down, the mine haul trucks will dump onto the ROM stockpile. A FEL (Front end loader) will be used to reclaim the ROM material and deliver the material to the dump pocket. The ROM stockpile will also be used to feed the crusher if the mining operations are suspended. Ore from the secondary screen feed conveyor will be transported to the secondary vibrating screen. Screen undersize material will be conveyed to the 3,000 ton heap leach feed stockpile. Lime will be added to the stockpile feed conveyor from the 200 ton lime silo by screw conveyor for pH control at a rate of 2.7 kg/t. Screen oversize material will be conveyed to the secondary cone crusher. The secondary cone crusher discharge and jaw crusher product combine on the secondary screen feed conveyor back to the secondary screen. The crushing circuit is designed to handle all ore types and comprises a primary jaw crusher, a secondary and an optional tertiary cone crushers operating in closed circuit with a final product screen. 17.3.2 Heap Leach Facility The heap leach facility (HLF) is designed to allow crushed ore stacking to a maximum height of approximately 24m (measured vertically over the liner system), which results in a design capacity of 5.6 Mt. The HLF comprises the following:
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Conventional, three stage lift (nominally 8 m per lift), free-draining heap over a gently sloping heap leach pad (HLP) along the axis of the ridgeline west of the ADR plant; The leach pad will be graded and constructed in a nominally balanced cut-and-fill manner using locally borrowed (within the heap boundary) rock for structural fill, supplemented as needed by mine waste including waste rock and, if available, thaw-stable soil for lining the pad subgrade before placement of the liner system; Permanent and interim perimeter diversion channels and berms to manage surface water flows; Perimeter access and ore haulage roads; Leach pad liner system will be constructed in the steps described below: o Graded subgrade to provide a non-puncturing surface for the geosynthetic liner; o Leak detection using horizontal wick drains to operate as large-scale lysimeters; o Reinforced clay liner (500mm); o Primary geomembrane liner, 1.5 mm thick linear low density polyethylene (LLDPE), bottom side aggressively textured, and o Overliner gravel (crushed ore), 700 mm thick, with drainage pipes, to protect the liner from ripping during initial ore stacking and to minimize the hydraulic head directly over the geomembrane; Gravity drainage from the leach pad to the pregnant tank at the ADR plant or (in the case of an upset) the events ponds in double-contained and buried pipes;
Figure 17.3: HLF plan
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Figure 17.4: HLF Cross section 17.3.3 ADR Gold Recovery Plant 17.3.3.1 Carbon Adsorption The carbon adsorption circuit consists of a train of five cascading carbon columns. The pregnant or gold- enriched solution will be pumped to the carbon adsorption circuit across a stationary trash screen for removal of any debris from the heap leach pad. The solution will flow counter- current to the movement of carbon from column 1 to column 5. The solution overflow from the final column will discharge onto a screen in order to recover any carbon. The barren solution, which at this stage has had most of the gold in solution adsorbed, will discharge from the final carbon column and will be pumped to the barren tank.
Cyanide solution, caustic solution, antiscalant and make-up water are added to the barren tank as needed. Barren solution will be heated to increase solution temperature by 8oC before being pumped back to the leach pad in order to maintain the thermal integrity of the heap leach pad. On average, 5 ton of loaded carbon from the first carbon column will be pumped to the acid wash and stripping circuits each day. The carbon-in the second column will be advanced to the first and the process will be continued down the train. The carbon from the fifth column will advance to the fifth column and then freshly reactivated carbon will be added.
17.3.3.2 Desorption and Gold Refining The loaded carbon will be transferred to the acid wash vessel and treated with 3% hydrochloric acid (HCl) solution to remove calcium, magnesium, sodium salts, silica, and fine iron particles. Organic foulants such as oils and fats are unaffected by the acid and will be removed after the stripping or elution step by thermal reactivation utilizing a kiln. The dilute acid solution will be pumped into the bottom of the acid wash vessel, exiting through the top of the vessel back to the dilute acid tank. At the conclusion of the acid wash cycle, a dilute caustic solution will be used to wash the carbon and neutralize the acidity.
A recessed impeller pump will transfer acid washed carbon from the acid wash tank into the strip or elution vessel. Carbon slurry will discharge directly into the top of the elution vessel. Under normal operation, only one elution will take place each day.
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17.3.3.3 Carbon Stripping (Elution) After acid washing, the loaded carbon will be stripped of the adsorbed gold using a modified ZADRA process. The strip vessel holds approximately 5.0 ton of carbon. During elution containing approximately 1 % sodium hydroxide and 0.1 % sodium cyanide, at a temperature of 140 °C and 450 kPa, will be circulated through the strip vessel. Solution exiting the top of the vessel will be cooled below its boiling point by the heat recovery heat exchanger. Heat from the outgoing pregnant solution will be transferred to the incoming cold barren solution. A diesel-powered boiler will be used as the primary solution heater to maintain the barren solution at 140°C. The cooled pregnant solution will flow by gravity to the electrowinning cells. At the conclusion of the strip cycle, the stripped carbon will be pumped to the carbon-regeneration circuit.
17.3.3.4 Carbon Regeneration The stripped carbon from the strip vessel will be pumped to the vibrating carbon-sizing screen. The kiln-feed screen doubles as a dewatering screen and a carbon-sizing screen, where fine carbon particles will be removed. Oversize carbon from the screen will discharge by gravity to the 7.5 ton carbon-regeneration kiln- feed hopper. Screen undersize carbon will drain into the carbon-fines tank and then be filtered and bagged for disposal. A 250 kg/h diesel-fired horizontal kiln will treat 5.0 ton of carbon per day at 650°C, equivalent to 100% regeneration of carbon. The regeneration-kiln discharge will be transferred to the carbon quench tank by gravity, cooled by fresh water or with carbon-fines water, prior to being pumped back into the CIC circuit.
To compensate for carbon losses by attrition, new carbon will be added to the carbon attrition tank.
New carbon and fresh water are mixed to break off any loose pieces of carbon prior to being combined with the reactivated carbon in the carbon holding tank.
17.3.3.5 Refining Pregnant solution will flow by gravity to a secure gold room. The solution will flow through one of two 3.54 m3 electrowinning cells. Gold will be plated onto knitted-mesh steel wool cathodes in the electrowinning cell. Loaded cathodes will be power washed to remove the gold-bearing sludge and any remaining steel wool. The gold-bearing sludge and steel wool will be filtered to remove excess moisture and then retorted to remove any mercury. The retort residue will be mixed with fluxes consisting of borax, silica and soda ash before being smelted in an induction furnace to produce gold doré and slag. The doré will be transported to an off-site refiner for further purification. Slag will be processed to remove prills for re-melting in the furnace. The gold bars will be stored in a vault located in the gold room prior to secure off-site transportation by aircraft.
17.3.3.5 Reagents Sodium cyanide briquettes will be delivered to site in containers and in 1 ton super sacks contained in a wood frame. The briquettes will be mixed in the cyanide mix tank and subsequently transferred to the cyanide solution storage tank. The concentrated cyanide solution will be added to the barren tank at a rate of 0.2 kg/t of ore. Cyanide will be used in the carbon strip circuit at a concentration of 0.1%.The principles and standards of practice for the transport to site and handling of cyanide on site will be in accordance with the guidelines set out in the International Cyanide Management Code (ICMC).
Sodium Hydroxide (caustic) will be supplied to site in 1 ton totes. The caustic will be mixed and stored for distribution to the acid wash and strip circuits. The caustic will be used to neutralize the acid in the acid wash circuit. A solution of 1.0% caustic will be mixed with barren solution in the carbon strip circuit.
Hydrochloric acid and antiscalant solutions will be supplied to site in 1 ton totes. The solutions will be metered directly from the totes for distribution in the plant.
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Hydrated lime will be delivered to the site in bulk by trucks and stored in a 200 ton lime silo. The lime will be delivered at a rate of 2.7 kg/t of ore by screw feeder onto the heap leach feed conveyor during heap loading operations.
17.3.3.6 Laboratory An assay and metallurgical laboratory will be equipped to perform sample preparation and assays by atomic absorption, fire assay, and CN soluble analyses. The facility will be equipped to prepare and analyze up to 3,600 samples per month. The laboratory facility will support exploration, mining, minor environmental sampling, TSS monitoring and processing. The majority of the environmental samples will be sent off-site to an accredited laboratory for third party reporting. The laboratory has space available for process optimization and test program.
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18 PROJECT INFRASTRUCTURE
The ATO property is situated in the eastern side of the Mongolia, which located in Tsagaan Ovoo soum of Dornod aimag. 660 km east of the Ulaanbaatar city, northwest of the Choibalsan city and 38 km east of Tsagaan Ovoo soum. The closest town, Tsagaan ovoo soum is located on shore of Hoovor nuur is medium developed infrastructure.
The both of Infrastructure and construction building are designed for Mongolian harsh weather. Following Construction facilities would build in the project site:
Process-related facilities Laboratory Warehouse and maintenance areas Office building Fuel storage Explosives storage Power supply Water supply Heap leach facilities and ponds Camp 18.1 Facilities
Following major buildings constructed at the ATO polymetallic mine project include: 18.1.1 Processing Plant Processing plant, which has dimension 28 m width x 50 m length x 20 m height, built by sandwich technology has triangle top. The building contains workshop, inventory side, gold room and management room.
Figure 18.1: Processing Plant 18.1.2 Laboratory Laboratory building locates close to Processing plant. Office building next to laboratory building. The both building are single storey, office building 400m2 to house some administrative and technical staffs.
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Figure 18.2: Laboratory building 18.1.3 Warehouse and maintenance areas Warehousing will be a 2 story building, which combination of multitask rooms and a warehousing area. Selected rooms will be air-conditioned/heated to protect sensitive equipment. Such as second story includes inventory storage, office room, restroom and electrical panel room. Another story to house separate truck wash pad with high pressure monitors and oil separator augments the maintenance facilities. Adjacent to the wash pad is a reinforced concrete pad for tire and large component maintenance work.
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Figure 18.3: Warehouse and maintenance area
18.1.4 Office building The office building is a single-story, 750 m2 modular building to house all administrative, and technical staffs. Building space is also allocated for meetings and training.
Figure 18.4: Office building
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Figure 18.5: Office plan 18.1.5 Fuel depot Diesel and gasoline are purchased in bulk and stored on site at 6 refueling depots. Both have been constructed with full containment systems in the event of tank rupture. Mining and on-site diesel powered mobile equipment are fueled at the storage tanks.
There are following fuel tanks will constructed at project site. All tanks are made by steel
250 m3 tank one piece 60 m3 tank four piece /umderground/ 25 m3 tank one piece 18.1.6 Explosive storage The Explosive storage is inside gravel dam which is located 3 km from open pit. All explosive, detonators and transfer wires saves in separate containers. 18.1.7 Camp The mining camp constructed at the ATO project which has capable for 300 staffs include:
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Kitchen capacity for 250 people. Building space is also allocated for restroom, cold storage area and staff room.
Laundry building Heating plant Septic system 18.2 Water Supply
RPS Aquaterra done all water exploration and found capable source of water 420 l/sec for 20 years. The Project uses water rights with a total consumptive of 11.02 litres per second.
Table 18.1: Water usage
Total usage No User l/sec м3/hour 1 Camp 0.6 2.16 2 Open pit 2.65 9.54 3 Processing plant 7.77 27.97 Total 11.02 39.67
Figure 18.6: Water supply plan 18.3 Power Supply
The project’s normal operating demand load is estimated to be 2.4 megawatts. The project will connected to Diesel Power generator /CAT 3512B/ which are generates 3.6 mWt.
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Table 18.2: Power usage
No User Eac unit power usage, кWт 1 Camp 425.67 2 Open pit 51.5 3 Processing plant 1,938 Total 2,415
18.4 Roads
The mine access road connects the project site to Choibalsan city 120 km, approximately 660 km from the Ulaanbaatar city. The road is constructed as a gravel embankment. 18.5 Communications
A communication system is on site to support internet, VOIP, and data communications necessary for daily operation of the mine, plant, and office. The mine site also has good cell phone coverage. Plant operators, survey crews, supervisors, and the mine contractor all have portable hand-held radios for operational communications. Closest town Tsagaan ovoo’s population is 3,800.
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19 MARKET STUDIES
Gold and silver production can be competitively sold to numerous reputable smelters and refiners throughout the world on a regular and predictable basis. Demand for gold and silver is presently high, with December 2016 gold and silver spot prices ranging from approximately $1,275 to $1,334 per ounce and $17.68 to $22.11 per ounce, respectively.
The large number of available markets for ATO allows gold and silver to be readily sold on the spot market. STEPPE expects that the production from ATO will be marketed through the contracts currently in place for producing mines.
The gold and silver prices used in the economic analyses are $1,306.6 per Au oz and $21.6 per Ag oz.
Table 19.1: LME Historical Gold and Silver price, (USD/oz)
Year 2009 2010 2011 2012 2013 2014 2015 2016 2017 average Gold 972.35 1224.53 1571.52 1668.98 1411.23 1266.4 1160.06 1250.74 1233.46 1,306.6 USD/oz Silver 14.67 20.19 35.11 31.14 23.79 19.07 15.68 17.14 17.38 21.6 USD/oz
Figure 19.1: LME historical Gold price chart
Figure 19.2: LME historical Silver price chart
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20 ENVIROMENTAL STUDIES, PERMITTING, and SOCIAL IMPACT
This chapter is prepared according to “Preliminary assessment of mineral mining potential, general requirements for feasibility study of mining projects and regulations for submitting feasibility study” issued by Ministry of Mineral resource and Energy on 17th of April, 2012. This assessment was prepared by CGM. STEPPE is preparing updates for current environmental issues based on the current regulations and expected to be completed and submitted by the Fourth Quarter of 2018. There is a reasonable expectation that all government approvals will be issued.
This chapter contains project implementer’s monitoring system on environmental protection, legal environment, protection plan, revision program and mine closure and rehabilitation plan which are estimated to have expenditure of 20.6 million tugrik (MNT).
Table 20.1: Details on the cost for environment
PARAMETER TOTAL PERCENTAGE Mine closure and rehabilitation plan 15,482.91 74.95 Technical rehabilitation 5,884.01 28.48
Biological rehabilitation 8,386.8 40.60 Mine closure 783.9 3.79 Examination 428.1 2.07 Environmental protection plan 5,175.8 25.05 Air quality 224.1 1.08 Soil, surface 1,778.2 8.61 Plants 951.1 4.60 Water 445.6 2.16 Animals 254.0 1.23 Findings of historical and cultural value 55.0 0.27 Public Collaboration 979.0 4.74 Others 488.7 2.37 Total 20,658.7 100
76.9% of the total environmental cost is for mine closure and restoration while 25.5% is estimated for environmental protection and revision work during mining process. 20.1 Monitoring system on environmental protection
STEPPE shall comply with MNS ISO 14001:2005 standard that provides requirements for the design, implementation and certification of an environmental management system and MNS ISO 14004:2005”, a “General Guidelines On Principles, Systems And Supporting Techniques environmental management system” when preparing guidance of monitoring system on environmental protection. Environmental management system shall include:
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• Human resource to identify duties and responsibilities • To organize seminars and training course and invite participants • Communication ( to receive, to document and to answer internal formal communications as well as contacts of interested foreign parts) • Documents and them • Revision control on activities • To supervise and do measuring • To determine whether project meet the standards and requirements • To control the records /to oversee whether the project meet both STEPPE and the country’s legislation and regulation on environment and their requirements; and to keep track of records of results. • Monitoring analysis 20.2 List of mining licenses and regulations to be applied
The project implementer shall comply with the relevant legislation and standards of Mongolia.
Table 20.2: List of environmental laws relevant to placer gold and mixed metal deposit mining project.
Date of Date of № Name of law enactment amendment 1 Minerals law 2006.07.08 2012.05.27 2 Law on air pollution payment 2010.06.24 2012.05.27 3 Law on air 2010.06.24 2012.05.27 4 Law on fauna 2000.05.05 2012.05.27
5 Environmental protection law 2012.05.17
6 Law on environmental impact assessment 2012.05.17 7 Law on natural resource use fee 2012.05.17
8 Law on land 2002.06.07 2012.05.17 9 Law on disaster protection 10 Law on subsoil 1988.11.29 1995.04.17 11 Law on land fee 1997.04.24 2005.07.01
12 Law on Prohibiting Mineral Extraction and Exploration in Headwaters of Rivers 2009.07.16 and Forest Resource Areas
13 Law on Soil Protection and Prevention of Desertification 2012.05.17 14 Law on special protected areas 1994.11.15 2008.12.19 15 Law on water 2004.04.22 2012.05.17
16 Law on water pollution fee 2012.05.17 17 Law on culture 1988.11.19 1995.04.17 18 Law on protecting cultural heritage 2001.06.08 2005.06.02
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Some standards relevant to environmental protection and rehabilitation work is listed below.
Table 20.3: List of standards
№ Standard Title 1 MNS 5914:2008 Environment. Land reclamation 2 MNS 5915:2008 Environment. Classification of land destroyed due to mining activities. 3 MNS 5916:2008 Environment. Requirements for fertile soil removing and its temporary storage during the earth excavation. 4 MNS 5917:2008 Environment. Reclamation of land destroyed due to mining activities. General technical requirements. 5 MNS 5918:2008 Environment. Re-vegetation of destroyed land. General technical requirements. 6 MNS 0900:2005 Environment. Health protection, Safety. Drinking water. Hygienic requirements and control. 7 MNS 4943:2011 Water quality. Waste water 8 MNS 4586:1998 Water quality. General requirements. 9 MNS 3597:1983 Environmental protection, Water space. Common requirement protection of surface and underground water from fertilizer pollution. 10 MNS 3342:1982 Environmental protection, Common requirement of protection of underground water from pollution. 11 MNS 5885:2008 Acceptable concentration of air pollutant elements.General Technical requirements. 12 MNS 3297:1991 Environmental protection. Soil. Soil sanitary index norm of populated area. 13 MNS 6148 2010 Water quality. Maximum limit of substance contaminating the ground water.
20.3 Environmental impact: identifying, reducing and eradicating plan
This section contains activities of the detection process of environmental impact which may arise from ATO activity. Environmental baseline research was made on the basis of geo-ecological research statement carried out by “Eco Trade LLC” made in “Urd Tsagaan Ovoo” exploration area with “6727X” mining license. Environmental adverse impacts are categorized as follows:
• Surface, soil • Plants • Surface and underground water • Fauna • Air • State and local special protected areas • Memorable pieces of historical and cultural value • Seminars for public on Environmental issue • Other issues 20.3.1 Soil layers Impacts on soil layers
• Artificial landshaft will be formed instead of natural landshaft of 164.2 hectare out of 362 hectare to be damaged by the project operations. Following objects will be included in the process. - 44.5 hectare mining winning with depth of 240-260 meter - 66.5 hectare outside waste dump - 33.2 hectare lean ore dump - 20 hectare oxidized ore dump will be formed in the surface. • Due to emergence of household and factory originated liquid or hard waste and the absence of measurements of collecting, eliminating and disinfecting them, it may have adverse impact on plants, soil fertility, physical and chemical parameters, surface and soil water, air quality as well as human and animal health.
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• Attempts to establish ditch or canal by outside waste dump and fertile soil dump in order to protect objects may flow and erode by ran water may cause surface erosion. • Project operations may cause healthy soil to be pressed down under the facilities and grind down owing to traffic. • In the framework of the project, 1,529,298 cubic meter fertile soil will be removed and due to wrong method of soil conservation, fertile soil may flow and to be lost by the rainwater, and then may form dust. • Adverse impacts from the project operations may include bunch of crossroads among mine facilities. • Transportation will be made through 260km dirt road from ATO, placer gold and composite metal deposit to Undurkhaan city of Khentii province and then 200km asphalt road from Undurkhaan to Baganuur. A soil may be grinded down while transporting through 260km improved dirt road and due to high dust rate, air quality may be affected. However, during the mining period, maximum 2 trucks a day, minimum 2 trucks in 10 days go through the road. Thus, impacts related to transportation is relatively low. Number of trucks to be used for concentrate transportation is demonstrated below.
Preventing and reducing future adverse impact on soil
• Landshaft formed by human activities will be restored to its approximate original contour and undergo technical rehabilitation according to MNS 5917:2008 standard-“Reclamation of land destroyed due to mining activities. General technical requirements” • Replanting work will be made in 44.5 hectare area equal to open pit mine area. • Concentrate will be transported through 260km dirt road from project area to Undurkhaan city and the related works to be made are: - to create landfill sites along the road - To locate traffic signs and do modification every year. - To establish flood control channel in parts of the road susceptible to water flooding. - To make plant baseline research along the road and to make supervision every year. • In the framework of ATO project, we will create centralized landfill site and make waste management program. • Temporary waste storage area will be created. • The area where the project is proposed to be implemented is a pasture land and issues of the pasture of the herders living around the area will be solved prior to beginning of the project implementation. 20.3.2 Plants Future adverse impacts on plants
214 plant species of 41 families were recorded in the project area and among them, 6 are rare plants while 2 are very rare plants. Thus, these plants may be pressed down and destroyed by facilities and other objects.
Table 20.4 Rare and very rare plants species recorded in project area
№ Scientific name Mongolian name Status Occurrence 1 Allium anisopodium Ledeb. Шувуун Сонгино Rare Rare 2 Allium odorum L. Анхил Сонгино Rare Rare 3 Iris potanini Maxim. Потанинийн Цахилдаг Rare Rare 4 Stellaria dichotoma L. Ацан Ажигана Rare Rare 5 Astragalus scaberrmus Bge. Ширүүн Хунчир Rare Rare 6 Glycyrrhiza uralensis Fisch. Урал Чихэр өвс Rare Occasional 7 Dasiphora parvifolia Fisch. ex Lehm. Жижигнавчит Боролзгоно Very rare Occasional 8 Vincetoxicum sibiricum L. Сибирь Ерөндгөнө Very rare Occasional
Preventive and protective measurements
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• To begin planting research and experiments since the first year of the project implementation and carry out biological rehabilitation under successful experiment. • If those rare plant species are to be pressed down by the facilities, plants will be installed in different location. • To collaborate with the locals in collecting the seeds of aboriginal plant species and use them in rehabilitation experiment and research work. • To ensure plant sustainability in outside waste dump areas. 20.3.3 Water Manual and drilled well waters are used in daily water use of animal and human and we will work in securing them every year. Location of water sites made during baseline research in 2011 is demonstrated in picture below.
Research on drinking water quality of the households living near the project area will be made not less than 1 time in 3 months. Location of the sites where inspection will be made are shown in the picture below while parameters for comparison are described in table below. Inspection sites of the projects will not be limited as sites shown the picture. 20.3.4 Air quality • We will work out in making clear on the articles on environment in the cooperation agreement with partner companies of the project and work under inspection management program of its implementation. • We will take possible measures to reduce fume and smoke from machines and equipment. • Air quality inspection program and climate measurement will be held every year. • Special study on household susceptible to adverse impacts of drilling, blasting, vehicle noises and thrilling will be made as well as to inform to herders, to hear their voices and to solve if possible. Always inform to near households before starting drilling or blasting. 20.3.5 Fauna As mentioned above, earth works will not be done in any area except 362 hectare area susceptible to erosion and we will perform to be environmental impacts as low as possible.
Measures will be taken against hunting and workers and people who may come to the area related to the project will be introduced with regulations about illegal hunting.
• To limit traffic near the sites where animals drink water and protect them in order to ensure their tranquility. • To protect aboriginal animals of the region. • To study animal species of the region through households living in the area for a long time.
Study on animal species, their population, abundance and living condition will be carried out not less than once a year. 20.3.6 Archeological findings As specified in Article 27.8 of Mongolian law on Protection of Cultural Heritages, which says that preliminary prospecting and research shall be carried out by professional paleontological, archaeological or ethnological scientific organizations for the assessment, prior to issuing land for purposes of economic activity associated with settlement, construction, paving new roads, establishing hydro power plants, conducting agriculture, mine prospecting and exploitation, Khurelsukh.S and Munkhbayar.L, scientists of Institute of Archaeology of
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Mongolian Science Academy, carried out detailed advanced exploration for two times, in 2011 and 2012 respectively.
7 out of 37 archeological artifacts which were discovered through the exploration are located in license area of the mining, however, are not expected to have direct impact from project activities.
Archaeological findings cannot be moved without authorization of the relevant administration. Thus, protective exclusive zone will be created around 7 archaeological sites.
In the case of finding something of historical and cultural value at the time of project implementation, a person shall comply with Article 38.3 “If tangible cultural heritage is discovered during the possession and use of subsoil, the subsoil user shall stop the work, and urgently inform the Governors of the soum and district, police, and the organizations responsible for the issue.” 20.3.7 Public collaboration • Company shall have specialist or department of environmental affairs. • To organize regular training course about environmental protection and to oversee for its implementation. • To ensure participation of environmental specialists in annual courses and seminars on environment and rehabilitation as well as to support exchange of experiments with other companies implemented rehabilitation work successfully. • To organize seminars for the locals about recent and further environmental programs to be carried out not less than once in a year. And seek their participation to the program • To support and finance local eco-friendly activities • To provide workers with handouts on environmental policy, regulations and standards. • As specified in Article 9.1 of the law on environmental impact assessment, environmental management plan shall be conducted by organization authorized to perform environmental impact assessment, be approved by relevant administration, and shall take measures in accordance with the program. • To purchase needed equipment used for permanent inspection work. 20.3.8 Other environmental issues 20.3.8.1 Uranium problem Initial exploration license holder COGEGOBI conducted an exploration for gold and uranium at the ATO and exploration was extended for Au and base metals by CGM after all as mentioned at the Section 6.1 of this report. In 2012, CGM requested to the Atomic Energy Agency of Mongolia (AEAM) to conduct an independent analysis at the ATO to confirm any radioactive mineral presence. AEAM examined based on the samples from specific observation points, 2 drill holes from a mineralized zone, drill pads and drinking water at the ATO, as well as drill cores from UB core house. Thus, official act with number of 11/124 dated on January 18, 2013 from AEAM was sent to CGM with confirmation of no radioactive elements were found.
20.3.8.2 Arsenic problem Arsenic (As) mostly derives from arsenopyrite and rarely exists at ATO. As will not present an environmental problem on the heap leaching process at ATO. 20.4 General approach to Management of Acid Rock Drainage
The intent of this waste handling plan is to ensure that the management of mining activities and the implementation of mine closure at the ATO and it will be conducted according to best practice methodologies to eliminate the potential for contamination of the surrounding soil and water resources from the generation of Acid Rock Drainage (ARD).
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This will be achieved through ensuring:
• A high level of understanding of the ARD characteristics and potential of the various rock types encountered during mining to allow effective operational identification and management of ARD hazards; • The appropriate design and construction of waste rock landforms and ore stockpiles; • ARD management controls will be integrated with mine planning and operational grade control; • ARD management and mitigation measures will be clearly communicated and understood by mine management, supervisors and operators; • Monitoring programs will be designed and implemented that allows for the performance of ARD controls to be measured, and corrective actions applied in a timely manner when monitoring indicates ineffective ARD control; • Installation of appropriate water management features to ensure water captured from waste rock landform does not interact with natural water resources; and • The effective development and implementation of mine reclamation and closure plans that ensure ARD potential is minimized including the use of backfilling and flooding of mine pits upon cessation of mining activities. 20.4.1 Waste Rock Management Plan STEPPE will implement a predictive ARD operational management plan based on a data developed from a comprehensive waste rock and ore type characterization program that will commence prior to mining. The characterization of waste rock and stockpiled ore during operations will be used to verify the predictive ARD model and will be accomplished using on-site testing methods combined with periodic laboratory analysis at a suitable facility.
Leco Sulfur Analysis will be used to assess the percentage of unoxidized sulfur in the test material. The value of the Leco Analyzer lies in the fact that it produces accurate results in a relatively short period of time. This would allow almost immediate knowledge of critical chemical values that the rock possesses and therefore allows suitable placement of the rock (to limit the potential for double handling of material). The use of onsite sulfur analysis represents Good International Industry Practice and provides a real time tool to validate ARD risk and confirm the appropriate management controls. 20.5 Neutralization and storage of chemical solution and poisonous substances
Natrium cyanide, Hydrochloric acid and Copper sulfate 20.5.1 Natrium cyanide Neutralization
Neutralization must be carried out prior to all cyanide used in industrial process go to waste storage facility. Cyanide will be extracted by double sulfur oxide.
- - CN + SO2 + O2 → OCN + H2SO4 (1) atmosphere pH 8 – 10
When the package damaged and leak, if deemed necessary, wear face mask, eye and face shed from dust, PVC gloves and chemical resistant booth. If it’s dangerous for respiration, wear full-face air purifying respirator.
Storage
• All chemical products must be stored in dry, clean environment with good air conditioning and labelled in accordance with the instructions on the safety sheet.
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• Do not store acids and alkalis together. • Food products and smoking is prohibited in premises where cyanide is handled. • Chemicals must not be stored together with flammable and combustible material. • Chemical stores must not have open floor drains.
Store all chemicals in cold, dry places with air conditioning and keep them away from acids, metals, and sun light and food products. Prevent chemicals from polluting and damage. Always be sure if they are leaking. 20.5.2 Hydrochloric acid Neutralization
When the package damaged and leak, if deemed necessary, wear face mask, eye and face shed from dust, PVC gloves and chemical resistant booth. If it’s dangerous for respiration, wear full-face air purifying respirator.
Storage
Store all chemicals in cold, dry places with air conditioning and keep them away from acids, metals, and sunlight and food products. Prevent chemicals from polluting and damage. Always be sure if they are leaking. 20.5.3 Copper sulfate Neutralization
When chemical leaks, workers must wear chemical resistant clothes by standing above wind. When hard material is spread, sweep them and transport to waste or put in proper package. If a little bit of liquid spilled, infiltrate them in sand or other similar material then transport to waste upon washed by water.
Storage
Store copper sulfate in cold, dry places with good air conditioning and keep them away from heat emitting or mitigating substance. 20.6 Preventing poisonous substance from seeping into soil
• Processing plant shall implement following works related to occupational safety. • To formulate regulation on occupational safety. • To grant patterns for all technological process of the industry and ensure their implementation • To formulate general guidelines of occupational safety and relevant guidelines for all occupation. • Educate workers for danger of fire. • To coordinate occupational safety regulations of ore processing plant with innovations. • To have workplace assessment conducted by professional company. • Workers will get medical examination once in 3 months. • To produce special plan for improving workplace condition. • To ensure proper use of chemicals under relevant regulations with low environmental impacts. • To meet the standards of safety and hygiene of industrial facility. • To provide workers with work wear and protective equipment preventing from unwanted situation of the workplace.
Most of the industrial processes are conducted inside and oxidized ore leaching and dressing area, pond and waste embankment of the primary ore will operate under special methods and international regulations.
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In heap leach facility, usually geomembranes such as high-density polyethylene, linear low-density polyethylene and polyvinyl chloride are used. We have chosen to use 1.5 mm thick LLDPE geomembrane in internal layer of the facility on the basis of its potential, its design as well as international experience.
Bore holes will be settled around basement foundation of the leach field in order to monitor water infiltration during its building process. Sampling of the monitoring bore holes will be made daily and identify infiltration of solution into ore leaching.
Basement system of pregnant solution pond comprises of 2 mm thick HDPE geomembrane which allows detection of loss and comprises of compacted soil which formed 300 mm thick layer, 1.5mm thick LLDPE liner and geonets.
Storm water pond foundation will be constructed with 2 mm thick HDPE liner which covering 300 mm thick compacted soil.
In the occasions of neither power goes off or pump is broken or strong wind and heavy rainfall, solution from pregnant solution pond may flow into near storm pond which consists of LLDPE liner. This storm water pond potential is estimated to hold highest rainfall of 104 mm for 24 hours.
In the case of over loading and heavy rainfall, pregnant solution pond will transfer its overflow to storm water pond.
Waste storage facility shall meet following requirements.
• To have sufficient storage capacity for waste arise during total mining period of the processing plant. • To have no adverse impact on environment as well as public safety. • To ensure sustainability both during mining period and closure. • To be located distant from rivers and water resources. • to be protected from possible danger of flooding • To be distant from settled or high populated areas. • To ensure and monitor condition that does not allow waste water infiltration into underground water • To have low construction cost of waste storage facility • To have low exploitation cost • To ensure minimum environmental impact relevant to waste transportation • To have low cost of closure
Inspection and monitoring of production process will be made for technological waste of the ore processing plant according to Mongolian and international standards.
There are two factors which have high possibility of affecting environment during exploitation period of the waste storage facility:
1. Small particles of waste accumulated in waste storage facility spread through surrounding environment and become dust when dry climate or wind occurs. 2. Waste water may leak from waste storage facility and infiltrate into soil and underground water.
Thus, these common issues must be solved in research and design planning level. There are several ways to prevent from dust and these are as shown below:
• To ram down by covering with poisonous chemical substances • To spray water • To layer with stone or soil • Or to layer with special material.
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20.7 Measures to take during special conditions
Mine emergency response plan shall be complied mine wide for special conditions. The mine manager shall develop and maintain a Mine Emergency Response Plan. The mine manager is responsible for its implementation. Mine emergency response plan outlines the response procedures and preventive measures that are essential for effective and timely management of an emergency situation. Prompt action and advance preparation can help save lives and protect financial investments in the event of emergencies such as mine fires, explosions or inundations.
Emergency response procedures help to organize and prepare personnel for emergency situations by establishing a common set of rules for training all emergency response personnel. The mine manager is responsible for providing training to all personnel involved in emergency operations. 20.8 Mining closure, rehabilitation, cost estimation
Total 362 hectare area is estimated to be damaged from project activities. 34% waste embankment, 18% outside waste dump, 12% open mine, 9% poor ore pile.
Figure 20.1: Damaged land from mine activities
The following operations are proposed to be held under the mining closure management plan of ATO gold- base metal mining project. 20.8.1 First phase activities of mining closure Following activities will be performed in the first phase of mining closure.
• To remove all mining facilities and equipment from the projected area • To build 2.5m tall lattice fence 5m away from upper bench of the mine in order to ensure animal and human safety according to guidelines for conducting technical and biological rehabilitation work in damaged land due to project operations and to locate signs on it.
Measures shall be taken for preventing animal and human to enter in the mining area and to get involved in accident. 20.8.2 Intense rehabilitation activity Following activities will be performed for mine rehabilitation program of ATO deposit mining project.
• Fertile soil removal, storage
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• Technical rehabilitation • Biological rehabilitation
Fertile soil removal and storage
362 hectare area is expected to be damaged during project implementation. Its 360 hectare fertile soil will be removed and stored according to MNS 5916:2008 standard.
Due to ecological situation of the soil layer, 5 types of soil are categorized in mining license area of the project. But 3 types of soil are spread in the area for mining planning and construction of mining facilities and the soil layer varies by their attributes.
Thickness of the removal is determined on the basis of morphological features of each type of soil and the removing process was made according to that measurement.
As a result of the removal of fertile soil with above mentioned thickness in 360 hectare area, 1,529,298 м3 soil will be formed. Here are the measures to take in storing fertile soil.
• During earthwork process, fertile soil pile will be removed and stored according to MNS 5916:2008, Standard for fertile soil removal and storage. • Prevention work should be conducted to prevent fertile soil pile from blow down, to be destroyed by flood, to be covered with salt, to be contaminated and to be pressed down by solid body, stone and construction material or waste. Fertile soil pile should be located and shaped to provide best condition for storage as well as piled according to wind direction with the height less than 5 meters. 30.6 hectare area is needed for up to 5 meter-tall pile. • If there is deemed necessary to store fertile soil for more than 2 years, windshield should be built and plant perennial grassy plants in the shaped upper surface and sides of the pile. • Permanent works for the pile include monitoring of its storage, protection as well as maintaining.
Technical rehabilitation
Technical rehabilitation work shall be carried out through several machineries such as CAT 962H wheel loader, CAT 773G off-highway truck, CAT 390CL excavator, and CAT 773G bulldozer and CAT D8T off-highway truck.
Biological rehabilitation
Following works shall be conducted within rehabilitation work of ATO mining project. Biological rehabilitation work is projected in 362.1 hectare area which had undergone erosion and its expenditure has estimated to be 7 billion tugriks (MNT).
Training courses or take guidance from specialists and professional organizations on successful biological rehabilitation and maintaining the area recently rehabilitated. To place covering upon improving and processing the fertile soil. To conduct biological soil rehabilitation experiment. Planting To consider on maintaining the area where planting was made.
Seminars on biological soil rehabilitation
Biological soil rehabilitation work should be conducted and monitored according to preprocessed project including relevant professional organizations and specialists (agronomist, botanist, biologist, etc.)
To place covering upon improving and processing the fertile soil
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Biological soil rehabilitation work shall begin by relocating fertile soil stored in pile upon all technical phases conducted and soil of the pile or area had stabilized and compacted.
Before directly using fertile soil in biological soil rehabilitation work, laboratory testing of soil fertility, physical and chemical property, soil composition and heavy metal inspection shall be made. When the soil does not meet requirements for proper condition of covering, measures to improve soil and use mineral fertilizer can be taken.
Biological soil rehabilitation testing
Experiments shall be made under following directions during ATO project implementation process and the successful experiments will be introduced.
• Experiments on soil 1. To use soil without improvement measures 2. To initiate planting and take different improvement measures with bio-pulp, peat, dung or fertilizer. • Experiments on plants 1. To perform planting by choosing plants that can adapt to the weather condition of the region such as perennial plants. 2. To collect aboriginal plant seeds with the help of local people and use collected seeds in planting. 3. To set up greenhouse and create seedlings for rehabilitation work. 4. To conduct rehabilitation work with seeds and seedlings and compare the results 5. To study rare and very rare plant species of the project area and collect their seeds for the rehabilitation testing
Planting work
There are some plants appropriate for using in planting work of the ATO project considering weather condition and environment of the projected area. They are: wheat grass, brome grass, Lucerne, melilot, pea, bush, oat, barley, soy, pea, rapes, etc. Also, growing aboriginal perennial plants considering appetite of the livestock live in area would be effective.
Maintaining the planted area
Perennial plant species should be planted in biological rehabilitation work and monocyclic plant species should be chosen as covering plant. 3 years of maintaining and protection is needed for perennial plant species. 20.8.3 Rehabilitation activity Due to nature and weather conditions of that year, rehabilitation process may have different results. Thus especially, rehabilitation result of the first year must be revised in the first hand.
Monitoring cost and additional cost or cost for possible situation must be calculated in total rehabilitation expenditure. Local community shall take the money upon receiving rehabilitated area from mining company. The project monitoring expenditure is calculated as Environment and Tourism Minister and Mineral Resources and Energy Minister's consolidated order enclosure No.А-132/112 dated May. 17th 2010.
Monitoring expenditure of the project is calculated at following rates
• Rate for calculating monitoring expenditure-5% • Coefficient-1
With this above calculation rate, monitoring expenditure of the rehabilitation work is calculated to 737.1 million tugriks. Monitoring is deemed necessary in following areas.
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• Open mine and its surrounding area • Outside soil pile • Oxidized ore pile • Poor ore pile • Pond and its surrounding area • Main road • Processing plant • Auto service • Combustive and lubricating material storage facility • Waste site • Healthy areas that weren’t eroded or contaminated
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21 CAPITAL AND OPERATING COSTS
Capital and Operating costs were estimated only for oxide ore mining and processing operation. 21.1 Capital Costs (Investment)
Investment will not be required since the open pit mining operation is contracted as a contractor. For Processing plant’s operation management will be contracted out. A total Investment of $ 19.59 million will be required for the whole project. Investment for ROM per ton will be $ 3.75 in ore. The investment amount includes VAT, customs duties, transportation costs and installation costs.
Table 21.1: Project investment summary
TOTAL INVESTMENT Pre-mining Engineering Plant overburden design drawings construction Camp and Working capital and ore stock Total, and permits and equipment’s infrastructures, (60 days need), preparation 1,000 USD expenses, cost, 1,000 USD 1,000 USD expenses, 1,000 USD 1,000 USD 1,000 USD 2017-III 500 500 2017-IV 500 1,500 2,000 2018-I 500 6,007 1,000 7,507 2018-II 500 3,604 1,000 5,104 2018-III 1,201 645 1,846 2018-IV 1,317 1,201 114 2,632 Total 1,317 2,000 12,015 4,145 114 19,591
Table 21.2: Investment per ton of ROM
Investment item USD/t (ROM) Pre-mining overburden and ore stock preparation expenses 0.25 Engineering design drawings and permits expenses 0.38 Plant construction and equipment’s cost 2.30 Camp and infrastructures 0.79 Working capital 0.02 Total cash outflow, 1,000 USD 3.75
Working capital
Current assets are estimated as $ 0.11 million.
Pre-mining overburden and ore stock preparation expenses
In the third quarter of 2018, earthworks will be commenced and 241.9 thousand tons of soil will be removed until 302.4 thousand tons of ore will be removed and the heap leach area will be prepared. The contractor’s price will be $ 2.42 per ton and the total cost will be $ 1.32 million.
Engineering design drawings and permits expenses
The cost of consulting and construction design and approval is based on price offered by local companies. The work is estimated to cost $ 0.5 million every three months from the third quarter of 2017, totaling $ 2.0 million.
Plant construction and equipment’s cost
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The investment for processing plant is based on the price offered by China's equipment manufacturers and suppliers. Total $ 12.01 million will be required, with Leach pad of $ 3.07 million or 26%, Adsorption- Desorption-Recovery (ADR) Plant $ 2.38 million or 20%, Carbon regeneration $ 1.57 million or 13% respectively. Contingency is 13%.
Table 21.3: Plant construction and equipment cost
PLANT CONSTRUCTION AND EQUIPMENT’S COST USD SHARE Crushing and Screening Plant 800,000 7% Stacking Facilities 861,691 7% Leach Pad 3,071,981 26% Adsorption-Desorption-Recovery (ADR) Plant 2,384,012 20% Heavy Mobile Equipment 528,720 4% Building 373,150 3% Carbon regeneration 1,539,152 13% Piping 131,140 1% Workshop 128,481 1% Laboratory 562,449 5% Machinery 51,250 0% Software, computer 13,750 0% Furniture’s 2,000 0% Contingency 1,567,166 13% Total 12,014,943 100%
Camp and infrastructures cost
The project will need investment a total of $ 4.14 million for camps and infrastructure costs. Billing calculations are shown in the table below.
Table 21.4: Camp and infrastructures cost
Cost Component USD share Camp 500,000 12% Explosives storage 225,408 5% Electrical system 729,167 18% Yard preparation 2,822 0% Sewage treatment 11,622 0% Fencing 113,647 3% Fuel storage 141,769 3% Water supply system 1,225,000 30% Heating system 250,000 6% Other 945,385 23% Total 4,144,821 100%
21.2 Operating Costs (Operational expenses)
Mining cost
The operation of the open pit will be operated by the contractor. The price of a contract from local contracting companies will be $ 2.42 per ton. The cost structure of contractor is as follows.
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Table 21.5: Mining contractor cost per ton of rock mass
Cost item USD/t Supplies 0.51 Hourly Labor 0.22 Salaried Labor 0.09 Equipment Operation 0.90 Miscellaneous 0.19 Contractors management fee 0.50 Total cost 2.42
The contractor’s fuel price $ 0.734 per liter for the price quotation, while the main contractor operators calculated on a monthly basis salary and other engineers and technicians on a monthly basis. Estimated salaries of local mine workers. Salary expense includes Social Insurance and incentives. The salaries of operators, except for basic workers, are estimated at $ 3.5 per hour.
Table 21.6: Salaried labor operating cost
2018 Number salary per 2019 2020 2021 2022 2023-I Job tittle of person/USD IY Total workers month Dollars/Year Mine manager 1 2,500 7,500 30,000 30,000 30,000 30,000 2,500 130,000 Foremen 2 1,500 9,000 36,000 36,000 36,000 36,000 3,000 156,000 Engineer 1 1,500 4,500 18,000 18,000 18,000 18,000 1,500 78,000 Environmental 1 1,500 4,500 18,000 18,000 18,000 18,000 1,500 78,000 Technician 3 1,000 9,000 36,000 36,000 36,000 36,000 3,000 156,000 Purchasing 1 800 2,400 9,600 9,600 9,600 9,600 800 41,600 Secretary 1 700 2,100 8,400 8,400 8,400 8,400 700 36,400 Total 10 39,000 156,000 156,000 156,000 156,000 13,000 676,000 Social insurance-13% 5,070 20,280 20,280 20,280 20,280 1,690 87,880 Total salary and Social insurance 44,070 176,280 176,280 176,280 176,280 14,690 763,880
Table 21.7: Hourly labor operating cost
2018 IY 2019 2020 2021 2022 2023-I Years Total project Dollars/Year Driller 4,050 14,850 14,850 14,850 14,850 1,238 64,688 Blaster 8,100 29,700 29,700 29,700 29,700 2,475 129,375 Excavator Operator 12,150 44,550 44,550 44,550 44,550 3,713 194,063 Truck Driver 16,200 74,250 74,250 74,250 74,250 6,188 319,388 Heavy Equipment Operator 20,250 74,250 74,250 74,250 74,250 6,188 323,438 Utility Operator 8,100 29,700 29,700 29,700 29,700 2,475 129,375 Mechanic 8,100 44,550 44,550 44,550 44,550 3,713 190,013 Electrician 4,050 14,850 14,850 14,850 14,850 1,238 64,688 Equipment Maintenance 12,150 59,400 59,400 59,400 59,400 4,950 254,700 Laborer 8,100 29,700 29,700 29,700 29,700 2,475 129,375 Total 101,250 415,800 415,800 415,800 415,800 34,650 1,799,100
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Processing cost
The management of the processing plant will be contracted by a professional company. The company's price quotation is $ 5 per ton of concentrates, including the following costs.
Table 21.8: Processing contractor cost per ton of ROM ore
Cost item USD/t (ROM) Labor 0.99 Reagent 1.36 Consumables 0.92 Water 0.11 Electric (diesel generator) 1.30 Fuel 0.06 Contractors management fee 0.24 Total processing cost 5.00
Chemicals
The cost of chemicals was calculated as a result of a technological test using unit cost of the local supply company.
Table 21.9: Usage of chemicals
Price Unit total Project Cost per № Chemicals $/ton consumption, consumption per Total cost, ton of ore, kg/t year, t 1,000USD USD 1 Sodium Cyanide , (NaCN) 1,796.67 0.40 2,226.3 3,999.9 0.72 2 Lime,(CaO) 137.38 2.70 15,027 2,064.5 0.37 3 Activated Carbon , (C) 1,965.51 0.0150 83 164.1 0.03 4 Hydrochloric Acid, (HCl) 312.91 0.150 835 261.2 0.05 5 Caustic Soda, (NaOH) 562.83 0.220 1,224 689.2 0.12 6 (NH3) 1,263.06 0.030 167 210.9 0.04 7 Anti-Scalant 7,370.39 0.00438 24.38 179.7 0.03 Total 19,587.8 7,569.4 1.36
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Table 21.10: Processing plant workers operation cost
Total salary Total salary Total Salary per month Social insurance, and social per month, Job position position each person USD 13%, USD insurance, USD USD/year PLANT GENERAL Plant Manager 2 2,041 4,082 531 55,347 Senior Metallurgist 2 1,551 3,102 403 42,064 Mechanical Engineer 2 1,551 3,102 403 42,064 Electrical Engineer 2 1,551 3,102 403 42,064 Safety Engineer 2 1,551 3,102 403 42,064 Safety officer 2 816 1,633 212 22,139 Plant Secretary 2 612 1,224 159 16,604 Total 14 19,347 2,515 262,344 OPERATIONS Shift Supervisors 5 1,428.6 7,143 929 96,857 Crusher / Unloading Operator 5 816.3 4,082 531 55,347 Leach Pad Operator 5 816.3 4,082 531 55,347 Elution operator 10 816.3 8,163 1,061 110,694 Reagent / Services Operator 10 816.3 8,163 1,061 110,694 Day & Relief Operators 10 734.7 7,347 955 99,624 Total 45 38,980 5,067 528,563 MAINTENANCE Mechanical Supervisor 3 1,428.6 4,286 557 58,114 Welder 3 653.1 1,959 255 26,567 Electrician 3 734.7 2,204 287 29,887 Shift electrician 3 653.1 1,959 255 26,567 Operator /generator/ 3 653.1 1,959 255 26,567 Total 15 12,367.3 1,608 167,701 LABORATORY Laboratory Supervisor 2 1,429 2,857 371 38,743 Laboratory Analysts 5 1,020 5,102 663 69,184 Metallurgical Technician 5 653 3,265 424 44,278 Samplers 5 531 2,653 345 35,976 Total 17 13,878 1,804 188,180 Grand total 91 84,571 10,994 1,146,789
General administrations
Total cost of general project management (G&A) is $ 4.18 million and $ 0.8 per ton. The main expenditure for general management is the salary cost and is calculated as a local standard.
Table 21.11: Head office staff salary
TOTAL TOTAL YEARLY MONTHLY POSITION HEAD COUNTS MONTHLY SALARY, INCLUDING SALARY USD SALARY USD SOCIAL INSURANCE USD Managing director 1 5,000 5,000 67,800 Chief executive officer 1 2,000 2,000 27,120 Chief operating officer 1 4,000 4,000 54,240 Procurement manager 1 2,500 2,500 33,900 Office manager 1 1,000 1,000 13,560 Driver 3 600 1,800 24,408 Service staff 2 400 800 10,848 Total 10 17,100 231,876
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Total Operational cost
The total cost of the project is $ 48.81 million of which the open pit mine operating costs $ 18.22 million, the total cost of the processing is $ 26.14 million and the general administration cost is $ 4.18 million. The total cost per ton of ore is $ 9.3 or $ 332.8 per ounce. (Table 21.12)
OPEX per ounce, $ 450.0 390.8 400.0 350.0 330.7 333.6 325.7 329.4 300.0 250.0 200.0 150.0 100.0 50.0 - 2019 2020 2021 2022 2023-I OPEX per ounce, $
Figure 21.1: OPEX per ounce, USD
OPEX per ton of ore, USD 12.0 9.9 10.0 10.1 10.0 9.2
8.0 6.2 6.0
4.0
2.0
- 2019 2020 2021 2022 2023
Figure 21.2: OPEX per ton of ore, USD
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Table 21.12: Project operational cost and general expenses
OPEX Ore Total per OPEX Rock Mining Mining ROM Ore General General License processing operational ton per Project Period mass, cost, cost, ore, processing administration administration fee, cost, costs, of ounce, stage Tx1,000 $/t $1,000 Tx1,000 cost, $/t cost, $/t cost, $ 1,000 $1,000 $1,000 $1,000 ore, $ $
Pre - 2018 544.2 2.42 - - - - 1,317 mining IY stage
Mining 2019 1,868.3 2.42 4,521 1,200.0 5.0 6,000 0.8 960 55 11,536 9.6 330.7 stage
Mining 2020 1,909.4 2.42 4,621 1,200.0 5.0 6,000 0.8 960 55 11,636 9.7 333.6 stage
Mining 2021 1,795.4 2.42 4,345 1,200.0 5.0 6,000 0.8 960 55 11,360 9.5 325.7 stage
Mining 2022 1,850.0 2.42 4,477 1,200.0 5.0 6,000 0.8 960 55 11,492 9.6 329.4 stage
2023- Mining 105.6 2.42 255 427.3 5.0 2,136 0.8 342 55 2,789 6.5 390.8 I stage
Total 8,072.9 2.42 18,220 5,227.3 5.0 26,136 0.8 4,182 275 48,812 9.3 332.8
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22 ECONOMIC ANALYSIS
22.1 Revenue
The gold production will begin in the first quarter of 2019 and will produce 146,669 ounces of gold and 672,518 ounces of silver during the project period. Based on market research, gold price was valued at $ 1306.6 and silver price was $ 21.6 per ounce. The total gross revenue of the project will be $ 206.16 million of which gold revenue is $ 191.64 million and silver revenue will be $ 14.53 million.
Gold produced, oz 40,000.00 34,883.56 34,884 34,884 34,884 35,000.00
30,000.00
25,000.00
20,000.00
15,000.00
10,000.00 7,135 5,000.00
- 2019 2020 2021 2022 2023-I
Figure 22.1: Gold produced, oz.
Silver produced, oz 180,000 160,754 160,754 160,754 160,000 135,033 140,000 120,000 100,000 80,000 55,224 60,000 40,000 20,000 - 2019 2020 2021 2022 2023-I
Figure 22.2: Silver produced, oz.
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Table 22.1: Revenue
Gold in Gold Silver Gold Silver Gold Silver Total Silver in Gold Silver Period ROM produced, produced, price, price, revenue, revenue, revenue, ROM ore, oz recovery recovery ore, oz oz oz $/oz $/oz $ 1,000 $ 1,000 $ 1,000
2019 49,834 337,583 70% 40% 34,883.56 135,033 1306.6 21.6 45,579 2,917 48,496 2020 49,834 401,884 70% 40% 34,884 160,754 1306.6 21.6 45,579 3,472 49,051 2021 49,834 401,884 70% 40% 34,884 160,754 1306.6 21.6 45,579 3,472 49,051 2022 49,834 401,884 70% 40% 34,884 160,754 1306.6 21.6 45,579 3,472 49,051 2023-I 10,193 138,059 70% 40% 7,135 55,224 1306.6 21.6 9,322 1,193 10,515 Total 209,527 1,681,295 70% 40% 146,669 672,518 1306.6 21.6 191,638 14,526 206,164
Selling cost
Selling cost is comprised of royalties and refining costs. According to Article 47.3.3 of the Minerals Law of Mongolia, gold royalties are estimated at 2.5% of sales revenues and 5% for silver royalties. The refinery cost was $ 6 per ounce of gold and $ 0.6 per ounce of silver. During the project period, the selling cost will be $ 6.8 million. 22.2 Cash flow analysis
Non-Discounted cash flow before tax, interests, and depreciations (non-discounted)
Total net revenue from the project is $ 206.16 million, EBIDTA is $ 150.55 million and total non-discounted net cash flow is $ 130.96 million.
Accumulated Cash flow (non-discounted) 140,000
120,000
100,000
80,000
60,000
40,000
20,000
- 2017-III 2017-IY 2018-I 2018-II 2018-III 2018-IY 2019 2020 2021 2022 2023-I (20,000)
(40,000)
Figure 22.3: Accumulated Cash flow (non-discounted), thousand USD
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Table 22.2: Non-Discounted cash flow before tax, interests, and depreciations (non-discounted)
OPERATIONAL CASHFLOW INVESTMENT CASH OUTFLOW
Pre-mining Engineering overburden design Plant Working and ore drawings construction capital Total Periodical Total Selling Net stock and and (60 days cash Cash revenue, cost, revenue, OPEX, EBIDTA, preparation permits equipment Camp and need), outflow, flow, Accumulated Project 1,000 1,000 1,000 1,000 1,000 expenses, expenses, cost, infrastructures, 1,000 1,000 1,000 Cash flow, development Period USD USD USD USD USD 1,000 USD 1,000 USD 1,000 USD 1,000 USD USD USD USD 1,000 USD stages
2017- Pre-mining III - - 500 500 (500) (500) stage
2017- Pre-mining IY - - 500 1,500 2,000 (2,000) (2,500) stage
2018- Pre-mining I - 500 6,007 1,000 7,507 (7,507) (10,007) stage
2018- Pre-mining II - 500 3,604 1,000 5,104 (5,104) (15,112) stage
2018- Pre-mining III - - 1,201 645 1,846 (1,846) (16,958) stage
2018- Pre-mining IY - - 1,317 1,201 114 2,632 (2,632) (19,591) stage
Operation 2019 48,496 1,576 46,920 11,536 35,384 - 35,384 15,793 stage
Operation 2020 49,051 1,619 47,432 11,636 35,797 - 35,797 51,590 stage
Operation 2021 49,051 1,619 47,432 11,360 36,072 - 36,072 87,662 stage
Operation 2022 49,051 1,619 47,432 11,492 35,940 - 35,940 123,602 stage
2023- Operation I 10,515 369 10,147 2,789 7,358 - 7,358 130,960 stage
Total 206,164 6,801 199,363 48,812 150,551 1,317 2,000 12,015 4,145 114 19,591 130,960
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22.3 Income statement and profit ratios
EBIDTA of the project is $ 150.55 million, net profit after tax is $ 105.09 million and profit margin is 51%.
Table 22.3: Income statement and profit ratios
thousand USD 2019 2020 2021 2022 2023-I Total
Sales revenue 48,496 49,051 49,051 49,051 10,515 206,164
Less variable costs 13,112 13,255 12,979 13,111 3,157 55,613
VARIABLE MARGIN 35,384 35,797 36,072 35,940 7,358 150,551
in % of sales revenue 73.0 73.0 73.5 73.3 70.0 73.0
Less fixed costs 2,337 2,337 2,337 2,337 2,337 11,686
OPERATIONAL MARGIN 33,046 33,459 33,735 33,603 5,021 138,865
in % of sales revenue 68.1 68.2 68.8 68.5 47.7 67.4
Interest on short-term deposits ------
Financial costs ------
GROSS PROFIT FROM OPERATIONS 33,046 33,459 33,735 33,603 5,021 138,865
in % of sales revenue 68.1 68.2 68.8 68.5 47.7 67.4
Extraordinary income ------
Extraordinary loss ------
Depreciation allowances ------
GROSS PROFIT 33,046 33,459 33,735 33,603 5,021 138,865
Investment allowances ------
TAXABLE PROFIT 33,046 33,459 33,735 33,603 5,021 138,865
Income (corporate) tax 8,074 8,177 8,246 8,213 1,068 33,779
NET PROFIT 24,972.36 25,282.01 25,488.89 25,389.96 3,953.15 105,086
in % of sales revenue 51.5 51.5 52.0 51.8 37.6 51.0
Dividends 0 0 0 0 0 -
RETAINED PROFIT 24,972 25,282 25,489 25,390 3,953 105,086
22.4 Discounted cash flow (after tax)
Discounted cash flow calculates as discount rate at 10% per year, depreciation for 5 years, and scrap value is 40% at the end of the project and income tax is estimated 10% on under 3 billion MNT revenue (1usd=2,400MNT) up to 25% on above 3 billion MNT (or $1.25 million) revenue or more.
DCF analysis was carried out by Comfar Expert-III software. Net present value is calculated by September 27, 2017. As a result of the calculation, the net present value of the project is $ 67.78 million, the repayment period is 2.81 years and the IRR is 118%.
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Table 22.4: Discounted cash flow (after tax)
thousand USD 2017 2018 2019 2020 2021 2022 2023 Scrap
TOTAL CASH INFLOW - - 48,496 49,051 49,051 49,051 10,515 7,905
Inflow operation - - 48,496 49,051 49,051 49,051 10,515 -
Other income ------7,905
TOTAL CASH OUTFLOW 2,500 16,977 21,659 21,437 21,215 21,329 3,865 -
Increase in fixed assets 2,500 16,977 ------
Increase in net working capital - - 473 5 (10) 5 (359) -
Operating costs - - 13,112 13,255 12,979 13,111 3,157 -
Marketing costs ------
Income (corporate) tax - - 8,074 8,177 8,246 8,213 1,068 -
NET CASH FLOW (2,500) (16,977) 26,836 27,614 27,836 27,722 6,650 7,905
CUMULATIVE NET CASH FLOW (2,500) (19,477) 7,359 34,973 62,810 90,532 97,182 105,086
Net present value (2,441) (15,070) 21,656 20,258 18,565 16,808 3,665 4,357
Discounted Cash Flow (2,441) (17,511) 4,145 24,404 42,968 59,776 63,442 67,799
NET PRESENT VALUE at 10% 67,799
INTERNAL RATE OF RETURN 118%
MODIFIED INTERNAL RATE OF RETURN 118%
NORMAL PAYBACK at 0% 2.73 years 2019
DYNAMIC PAYBACK at 10% 2.81 years 2019
NPV RATIO 3.831497
Net present values discounted to Sep-17
Discounted cash flow (after tax), thousand USD 80,000 67,799 70,000 63,442 59,776 60,000
50,000 42,968 40,000
30,000 24,404 20,000
10,000 4,145 - 2017 2018 2019 2020 2021 2022 2023 Scrap (10,000) (2,441) (20,000) (17,511) (30,000)
Figure 22.4: Discounted cash flow (after tax), thousand USD
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Sensitiviy analysis of IRR 150%
140%
130%
120%
110%
100%
90%
80% -20% -16% -12% -8% -4% 0% 4% 8% 12% 16% 20%
Sales revenue Increase in fixed assets Operating costs
Figure 22.5: Sensitivity analysis of IRR
Table 22.5: Sensitivity of IRR
Variation (%) Sales revenue Increase in fixed assets Operating costs
-20% 87% 144% 125%
-16% 93% 138% 124%
-12% 99% 132% 122%
-8% 106% 127% 121%
-4% 112% 122% 119%
0% 118% 118% 118%
4% 123% 113% 116%
8% 129% 109% 114%
12% 135% 106% 113%
16% 141% 102% 111%
20% 146% 99% 109%
If the Revenue decreased by 20%, IRR 87% and if Investment increased 20%, IRR 99%. The IRR is 109% while operating cost increased by 20%. 22.5 Stream Analysis
The economic analysis (Tables 22.1 and 22.4) does not consider the impacts of the obligations of the STEPPE under the Stream Agreement to sell a certain amount of the gold and silver produced from the ATO Project. The obligations of the STEPPE under the Stream Agreement could make the actual economics of the ATO Project materially less favorable than the results of the economic analysis appearing in this report and could materially reduce the estimates for ATO Project revenue, net profit, net present value and internal rate of return.
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Table 22.6 indicates revenue net of stream. 2017 and 2018, there is $11.5 Million of stream deposit for each year and then production years indicated. The ounces produced were divided into two revenue groups, revenue group one is 25% of the gold ounces and 50% of the silver ounces produced times 30% of the metal prices used in this report to get the streamer sales and group two revenue is the remaining produced ounces times the metal prices. These are added together to obtain the total revenue.
Table 22.6: The Cash Revenue
The cash revenue in production years shown in the Table 22.7 and add to it the recognized/earned revenue from the upfront metal purchase price paid by the streamer in 2017 and 2018 to calculate a proper recognized revenue for each production year.
Table 22.7: Profit with stream
Discounted cash flow with stream calculated in Table 22.8 and graph shown in Figure 22.5.
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Table 22.8: Discounted Cash Flow with stream
Discounted cash flow 5% with stream, thousand USD 90,000
80,000 76,966 70,921 70,000 67,360
60,000 52,482 50,000
40,000 36,765
30,000 20,458 20,000 9,222 10,000 3,878
- 2017 2018 2019 2020 2021 2022 2023 Scrap
Figure 22.6: Discounted cash flow with stream, thousand USD
Adjustments to subtract fees associated with the stream from funds received and pay taxes on the purchase net of equity raise finds were calculated and shown in Table 22.9.
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Table 22.9: Adjusted Discounted Cash Flow with stream
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23 ADJACENT PROPERTIES
There are no adjacent properties with similar publically well-known mineralization to provide comparative mineralization characteristics in ATO project area.
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24 OTHER RELEVANT DATA AND INFORMATION
This technical report will support an updated Feasibility Study report and will source any update of Mineral Resource and Mineral Reserve report under Mongolian jurisdiction if necessary.
The QPs and GLOGEX are not aware of any other relevant data.
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25 INTERPRETATION AND CONCLUSIONS
25.1 Conclusions
In the opinion of the QPs, the ATO Project that is outlined in this report has met its objectives. Mineral Resources and Mineral Reserves have been estimated and a mine plan for oxide material was developed. The financial analysis based on the mine plan indicates the project that is viable and has a 5 year mine life for oxide ore. This indicates the original project objective of identifying mineralization that could support mining operations has been achieved. Work completed to date supports the project proceeding to detailed engineering leading to construction of the Project.
Mineral Resource was estimated based on the 32,791 raw samples from 238 diamond drill holes. Total length of the samples was 44,284.2 m. A total of 226 bulk density data was involved for the estimation. Grade shells for all three pipes were created based on the 0.1 g/t Au grade outlines. Validated topo surface, newly created quaternary sediment, oxide and transition surfaces were applied to tag rock codes for the estimate. Exploratory data analysis and spatial data analysis were performed on the raw and composite data. Grade capping was done for all metals and validated. Resource block model was created by 5 m x 5 m x 2.5 m blocks. Grade block models for Au, Ag, Pb, Zn and AuEq were estimated by OK and ID interpolations. Mineral Resource classification block model was tagged using interpolation parameters used for grade estimation with three passes. The resultant block models were validated by swath plots and global mean statistics against composite samples, grade block data of OK and ID results. OK block model was the initial grade estimation method to use resource estimate.
The QP estimates the ATO Mineral Reserve totals 5.23 Mt at an average of 1.25 g/t Au containing 210 thousand ounces of Au and at an average of 10.0 g/t Ag containing 1,681 thousand ounces of Ag within oxide pit; and combined Measured and Indicated Mineral Resources in addition to Mineral Reserves total 12.23 Mt at an average of 1.49 g/t Au containing 587 thousand ounces of Au, at an average of 9.99 g/t Ag containing 3,927 thousand ounces of Ag, at an average of 0.75 % Pb containing 202 million pound of Pb and at an average of 1.34 % Zn containing 362 million pound of Zn for fresh material. Estimated Mineral Reserves and Mineral Resources are based on a cut-off grade of 0.3 g/t AuEq for oxide and a cut-off grade of 1.1 g/t AuEq for fresh materials.
In order to comply with the CIM Definitions requirement of “reasonable prospects for eventual economic extraction”, a preliminary Whittle pit shells and ultimate pit designs for Pipe 1, Pipe 2 and Pipe 4 were used to constrain the Mineral Resource estimate using process recovery of 70% for Au and 40% for Au, a gold price of $1,306.6 and a silver price of $21.6, and other assumed pit parameters. Only blocks located within the pit designs are reported in the Mineral Reserve estimate.
The ATO deposit can be treated with heap leach methods. The HLP is designed to process oxide ore from three open pits. The proposed ATO process plant design will be based on gold heap leach technology, which will consist of crushing, stacking, and heap leaching with cyanide solution, adsorption from the pregnant solution onto carbon, carbon elution, and smelting.
Financial analysis of the ATO has determined that the Project will be economically viable and profitable. The target ore production rate during the life of the mine is 1.2 Mt per year.
Discounted cash flow calculates as discount rate at 10% per year, depreciation for 5 years, and scrap value is 40% at the end of the project and income tax is estimated 10% up to 25% on above $1.25 million revenue or more. Net present value of the project is $ 67.78 million, the repayment period is 2.81 years and the IRR is 118%.
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25.2 Risk Assessment
A summary of the project risks at this stage include:
Resource – Resource infill drilling may result in changes to the mineral resources, both in negative and positive.
Recovery – Although the expected performance of the leach pads and gold recovery appear reasonable given current experience, knowledge and previous test work, there still exist elements which have no specific test work to address. Ongoing metallurgical test work will be needed.
STEPPE’s current operation plan is to continue leaching the existing ore on the heap leach pad at the rate that it will accept solution. There is a risk that the current permeability cannot be maintained and that could cause the time for recovery of these ounces will be extended.
Rock Density - Based on an inadequate density data, a density block model was not interpolated. There could be a risk to reflect the resource number.
Gold Price - The issuer has no plans to hedge the gold production from the ATO Project. Therefore, the project will be fully exposed to the risk and reward associated with the market fluctuations in the price of gold.
Financing and Liquidity - In order to develop and operate the project the issuer will have to raise the required financing. There is risk associated with an inability to raise the funds necessary to upgrade and continue the ATO operation.
Predicted NPV – The predicted NPV is a direct function of most of the factors included in the feasibility study. The aggregate risk to the NPV is deemed to be low as there is not a compendium of low risk factors, or of medium to high risk factors.
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26 RECOMMENDATIONS
The ATO Project shows positive economic potential, and should continue in operation with a new mining in the end of 2018.
The QP recommends the following:
• Assess the block model to determine if there are regions within the resource area limits, currently defined as inferred resource owing to the influence of Au assays that could potentially be upgraded for gold grade and confidence levels by additional drilling. Positive upgrades may result in addition to the measured and indicated resources, if areas of significance are identified. Especially Pipe 2 needs resource infill drilling. • To improve future modeling efforts, the Author QP recommends that the lithology type be less and specific for mining needs. Current lithology database has 24 different lithology codes. Some of those lithology types were similar and combined based on the available information ended up with 10 lithology types for further analyses in this report. • Near pit exploration drilling where mineralization is open along the trend to add resource to the model. Recommendation applies to especially Pit 4. • Continuing work on refinement of bulk density estimate of ore material. • Conduct a significant check assay program utilizing a number of labs in addition to ALS. While standard and blank analysis results of CGM QAQC program do not raise questions relative to ALS’s results, the check assay program ensures a test of any potential bias in the existing results. This recommendation applies to production sampling as well as for exploration data. • Additional metallurgical test work was needed for both Pit 1, Pit 2 and Pit 4 ore material. This will include bulk samples in large columns investigating crush size, agglomeration and leach rates as the mine advances.
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27 REFERENCES
2016 Annual Information Form, Centerra Gold Inc. SEDAR, March 31, 2017.
Altan Tsagaan Ovoo (ATO) Pilot Plant, Xstrata Process Support, 09 November 2012.
CGM Annual Exploration Report 2010, May 21, 2010
CGM Annual Exploration Report 2011, May 22, 2011
CGM Annual Exploration Report 2012, September 5, 2013
CGM Annual Exploration Report 2013, February 16, 2014
CGM Annual Exploration Report 2014, January 13, 2015
Estimate of mineral resources, potential mineral reserves and preliminary production and economic assesment of the ATO project, Eastern Mongolia, Centerra Gold Inc, January 19, 2012.
Findings from Groundwater Exploration works undertaken in the Altan Tsagaan Ovoo area (Estimated Exploitable Groundwater Reserves) for supply to the ATO Mine Prospect, by RPS Aquaterra, August 2012.
High-Level Legal Due Diligence Report relating to the mineral licenses of Centerra Gold Mongolia LLC for STEPPE Gold LLC, by ELC LLP Advocates, May 25, 2017.
Khurelsukh S., Munkhbayar L., 2011, Detailed archeological exploration report in the exploration license 6727X in Tsagaan-Ovoo, Dornod province for Centerra Gold Mongolia.
Large Scale Column Test Report by Boroo Gold at Boroo Mill Plant, 2012.
L.Baasandolgor et al., Report on exploration work conducted at ATO precious and base metal deposit in Altan Tsagaan Ovoo area in 2010-2011, Ulaanbaatar 2012.
Preliminary Assessment of Oxide gold ore composites from the Altan Tsagaan Ovoo (ATO) gold project, ALS Metallurgy, G&T Metallurgical services, May 15, 2012
T.Di Feo, L. Kormos, D. Fragomeni, M. Hoffman, S. Wojtowicz and Y. Boudreau, Mineralogical and Metallurgical Test Program for Centerra Gold Inc., Xstrata Process Support, June 30, 2011.
T.R. Raponi, 2012, Metallurgical Testing and Preliminary Process Design, Centerra Gold Inc., Toronto, Canada.
Social Baseline Assessment, Altan Tsagaan Ovoo, Dornod Aimag, Mongolia by AATA International Inc with association of Eco Trade LLC, Denver, Colorado, USA, February 2012.
Victor Vdovin, October 17, 2011, Geotechnical Assessment of ATO slopes, Centerra Gold Inc, Toronto, Canada.
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