mara rosa project, goiás state, brazil (latitude …av. afonso pena, 4001 12º andar - bairro serra...

Av. Afonso Pena, 4001 12º andar - Bairro Serra Belo Horizonte - MG - Brasil - CEP 30.130-009

Mara Rosa Project, Goiás State, Brazil (Latitude 13°58.395′ S, Longitude 49°10.690′ W)

Pre-Feasibility Study

Prepared by Coffey Consultoria e Serviços Ltda on behalf of:

Amarillo Gold Corporation

Effective Date: 28 October 2011

Qualified Person: G. Keith Whitehouse - BSc (Geology and Geography), MAusIMM (CP)

Qualified Person: Chris Witt - BSc, DipMet, MAusIMM (CP)

Qualified Person: João Augusto Hilário - BSc (Min Eng), MAIG

Qualified Person: Clive Saunders – BSc, CGeol FGS,TMIE Aust, M.Zwe.I.E

Qualified Person: Frank Baker – BMet, MMet, MIMMM, MAusIMM

Qualified Person: Norman Lock - BSc, PhD, CGeol FGS, PrSciNat.

220810

Coffey Consultoria e Serviços Ltda

Mara Rosa Project, Goiás State, Brazil – 220810 Pre-Feasibility Study – 28 October 2011

Author(s): G. Keith Whitehouse Geologist BSc (Geology and Geography), MAusIMM (CP)

Chris Witt Associate Consultant - Metallurgy BSc, DipMet, MAusIMM (CP)

João Augusto Hilário Manager Technical Services BSc (Min Eng), MAIG

Clive Saunders Principal Tailings Engineer BSc, CGeol FGS,TMIE Aust, M.Zwe.I.E

Frank Baker Project Manager, Amariloo BMet, MMet, MIMMM, MAusIMM

Norman Lock Manager Geology BSc, PhD, CGeol FGS, PrSciNat.

Date: 28 October 2011

Project Number: 220810

Version / Status: Rev02.v01 / Final

Path & File Name: \\missfs01\data$\MINE\Operations\Projects & Proposals\Projects\Amarillo\03022_Mara_Rosa_PFS\Report Preparation\Supporting Reports\NI43-101\br_220810_PFS_Mara_Rosa_Amarillo.REV01.draft.docx

Print Date: Tuesday, 10 January 2012

Copies: Amarillo Gold Corporation (2)

Coffey Mining – Belo Horizonte (1)

Document Change Control

Version Description (section(s) amended) Author(s) Date

Rev01.v2 Peer review CWC 25/10/2011

Rev01.v3 Peer review CWC 29/12/2011

Rev02 Amarillo comments FB 6/1/2012

Document Review and Sign Off

(signed) “Norman P Lock”

(signed) “Curtis W Clarke”

Primary Author Norman Lock

Supervising Principal Curtis Clarke



Date and Signature Page

This report titled “Mara Rosa Project, Goiás State, Brazil, Pre-Feasibility Study” with an effective date of

28 October 2011 was prepared on behalf of Amarillo Gold Corporation by Hugo Hoogvliet, Keith

Whitehouse, Chris Witt, João Augusto Hilário, Clive Saunders, Frank Baker and Norman Lock and

signed:

Dated at Subiaco, Australia, this 11 day of January, 2012

(signed) “Gregory Keith Whitehouse”

Gregory Keith Whitehouse, BSc (Geology and Geography), MAusIMM (CP).

Geologist, Australian Exploration Field Services

Dated at Perth, Australia, this 11 day of January, 2012

(signed) ”Chris Witt“

Chris Witt, BSc (Chemistry), DipMet, MAusIMM (CP).

Associate Consultant – Metallurgy, Coffey Mining

Dated at Belo Horizonte, Brazil, this 11 day of January, 2012

(signed) “João Augusto Hilário”

João Augusto Hilário, BSc (MinEng), MAIG.

Manager Technical Services, Coffey Consultoria e Serviços Ltda

Dated at Perth, Australia, this 11 day of January, 2012

(signed) “Clive Thomas Saunders”

Clive Thomas Saunders, Dip. C.Eng. (Struct), CGeol FGS, TMIE Aust, M.Zwe.I.E.

Principal Tailings Engineer, Coffey Mining



Dated at Belo Horizonte, Brazil, this 11 day of January, 2012

(signed) “Frank Richard Baker”

Frank Richard Baker, BMet, MMet, MIOM3, MAusIMM.

Project Manager, Amarillo Gold Corporation.

Dated at Toronto, Canada, this 11 day of January, 2012

(signed) “Norman Philip Lock”

Norman Philip Lock, BSc, PhD, CGeol FGS, PrSciNat.

Manager Geology, Coffey Mining



Table of Contents

1 Summary ....................................................................................................................................... 1

1.1 Property Description and Ownership ................................................................................ 2

1.2 Geology and Mineralization ............................................................................................... 2

1.3 Status of Exploration ......................................................................................................... 3

1.4 Metallurgical Testwork ....................................................................................................... 3

1.5 Mineral Resource Estimates ............................................................................................. 5

1.6 Mineral Reserve Estimates ............................................................................................... 6

1.7 Mining Methods ................................................................................................................. 7

1.7.1 Geotechnical ........................................................................................................... 7

1.7.2 Mining Study............................................................................................................ 7

1.7.3 TSF.......................................................................................................................... 8

1.8 Recovery Methods .......................................................................................................... 10

1.9 Project Infrastructure ....................................................................................................... 12

1.10 Market Studies and Contracts ......................................................................................... 12

1.11 Environmental Studies, Permitting, and Social and Community Impact ......................... 12

1.11.1 Physical and Natural Environment ........................................................................ 12

1.11.2 Hydrology and Hydrogeology ................................................................................ 13

1.11.3 Social Environment ............................................................................................... 14

1.11.4 Waste and Tailings Disposal, Monitoring and Water Management....................... 14

1.11.5 Permitting .............................................................................................................. 15

1.11.6 Mine Closure ......................................................................................................... 15

1.12 Capital and Operating Costs ........................................................................................... 15

1.13 Economic Analysis .......................................................................................................... 18

1.14 Conclusions and Recommendations ............................................................................... 19

2 Introduction ................................................................................................................................ 21

2.1 Terms of Reference ......................................................................................................... 21

2.2 Qualified Persons ............................................................................................................ 21

2.3 Site Visits and Scope of Personal Inspection .................................................................. 23

2.4 Effective Dates ................................................................................................................ 23

2.5 Information Sources and References .............................................................................. 24

2.6 Units of Measure ............................................................................................................. 24

2.7 Previous Technical Reports ............................................................................................ 25

3 Reliance on Other Experts ........................................................................................................ 26

4 Property Description and Location .......................................................................................... 27

4.1 General Description ......................................................................................................... 27

4.2 Establishing Mineral Rights in Brazil ............................................................................... 29

4.3 Royalties and other agreements on the property ............................................................ 32



4.4 Environmental Liabilities .................................................................................................. 32

4.5 Permits required for Development .................................................................................. 32

4.6 Other Factors or Risks affecting Access, Title or ability to Work .................................... 32

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ....................... 33

5.1 Accessibility ..................................................................................................................... 33

5.2 Climate ............................................................................................................................ 33

5.3 Local Resources .............................................................................................................. 33

5.4 Infrastructure ................................................................................................................... 33

5.5 Physiography ................................................................................................................... 35

6 History ......................................................................................................................................... 36

6.1 Exploration History .......................................................................................................... 37

6.2 Metallica Exploration ....................................................................................................... 38

6.3 Amarillo Exploration ........................................................................................................ 39

6.3.1 Validation of Drill Hole Locations ........................................................................... 40

6.3.2 Waste Dump Volume ............................................................................................ 40

6.3.3 Surface Trenching ................................................................................................. 40

6.4 Historical Drilling .............................................................................................................. 40

6.5 Historical Resource Estimates ........................................................................................ 43

6.5.1 WMC Grade Tonnage Estimate ............................................................................ 43

6.5.2 Barrack Grade Tonnage Estimate ......................................................................... 43

6.5.3 Metallica Grade Tonnage Estimate ....................................................................... 44

6.5.4 Amarillo; CCIC Resource Estimate ....................................................................... 44

6.5.5 Amarillo; HCS & AEFS Resource Estimate 2010.................................................. 45

6.6 Historical Production ....................................................................................................... 46

7 Geological Setting and Mineralisation ..................................................................................... 47

7.1 Regional Geology ............................................................................................................ 47

7.2 Local Geology ................................................................................................................. 50

7.3 Property Geology ............................................................................................................ 52

7.4 Mineralisation .................................................................................................................. 55

8 Deposit Types ............................................................................................................................. 57

9 Exploration.................................................................................................................................. 58

10 Drilling ......................................................................................................................................... 59

10.1 Drill Hole Planning ........................................................................................................... 59

10.2 Technical and Support Staff ............................................................................................ 60

10.3 Drill Hole Setup ............................................................................................................... 60

10.4 Drilling Execution ............................................................................................................. 61

10.4.1 Down Hole Surveys ............................................................................................... 61

10.4.2 Driller’s Field Records ........................................................................................... 61

10.4.3 Core mark up (Field) ............................................................................................. 61

10.4.4 Collar Preservation ................................................................................................ 62



10.4.5 Core Shed Procedures .......................................................................................... 62

10.5 Densities .......................................................................................................................... 62

10.6 Sampling Method and Approach ..................................................................................... 63

10.7 Geological Data Collection .............................................................................................. 63

10.8 Opinion ............................................................................................................................ 64

11 Sample Preparation, Analyses and Security ........................................................................... 65

12 Data Verification ......................................................................................................................... 68

12.1 Data entry ........................................................................................................................ 68

12.2 Amarillo QAQC ................................................................................................................ 68

12.2.1 QAQC Results ....................................................................................................... 68

12.2.2 Due Diligence QAQC ............................................................................................ 69

12.3 Drillhole coordinates ........................................................................................................ 70

12.4 Topographic survey data ................................................................................................. 70

12.5 Downhole Survey ............................................................................................................ 71

12.6 Improvement of Drilling Programs ................................................................................... 72

12.6.1 Drill Rig Setup ....................................................................................................... 72

12.6.2 Down Hole Surveys ............................................................................................... 72

12.6.3 Data base .............................................................................................................. 72

12.7 Drill Program Assessment ............................................................................................... 73

13 Mineral Processing and Metallurgical Testing ........................................................................ 74

13.1 Introduction ...................................................................................................................... 74

13.2 Background ..................................................................................................................... 74

13.3 Metallurgical Domaining .................................................................................................. 78

13.4 Sample Selection and Head Grade Analysis .................................................................. 79

13.5 Testwork Programme ...................................................................................................... 79

13.6 Comminution Testwork Results and Interpretation ......................................................... 80

13.7 Metallurgical Testwork Results and Interpretation .......................................................... 82

13.7.1 Grind Size.............................................................................................................. 84

13.7.2 Lime Demand and Effect of pH ............................................................................. 85

13.7.3 Pre-Oxidation and Cyanidation Time Leach.......................................................... 87

13.7.4 Cyanidation Leach Residues................................................................................. 87

13.8 Geochemistry .................................................................................................................. 87

13.8.1 Introduction............................................................................................................ 87

13.8.2 Samples ................................................................................................................ 88

13.8.3 Testwork Programme ............................................................................................ 88

13.8.4 Acid-Base Chemistry ............................................................................................. 88

13.8.5 Multi Element Analysis .......................................................................................... 91

13.8.6 Conclusions ........................................................................................................... 92

13.9 Ancillary Testwork ........................................................................................................... 93

13.9.1 Site Water.............................................................................................................. 93

13.9.2 Settling / Thickening .............................................................................................. 93



13.9.3 Pulp Viscosity ........................................................................................................ 93

13.10 Key Design Criteria ......................................................................................................... 94

13.11 Process Flowsheet .......................................................................................................... 94

13.12 Further Testwork Recommendations .............................................................................. 95

14 Mineral Resource Estimates ..................................................................................................... 96

14.1 Data Utilised .................................................................................................................... 96

14.2 The Modelling Process .................................................................................................... 96

14.3 Model Classification ...................................................................................................... 102

14.4 Resource ....................................................................................................................... 103

15 Mineral Reserve Estimates ...................................................................................................... 106

16 Mining Methods ........................................................................................................................ 109

16.1 Geotechnical ................................................................................................................. 109

16.1.1 Previous studies .................................................................................................. 109

16.1.2 Deposit Geology .................................................................................................. 111

16.1.3 Geotechnical Data ............................................................................................... 112

16.1.4 Geotechnical Model............................................................................................. 114

16.1.5 Slope Stability Assessment ................................................................................. 117

16.1.6 Pre-feasibility conclusions ................................................................................... 120

16.2 Mining Study .................................................................................................................. 121

16.2.1 Pit Optimisation ................................................................................................... 121

16.2.2 Pit Design ............................................................................................................ 127

16.2.3 Mine Scheduling .................................................................................................. 130

16.2.4 Waste Rock, Ore and Low Grade Stockpiles ...................................................... 135

16.2.5 Tailings Storage Facility ...................................................................................... 141

16.2.6 Mine Production and Operating Parameters ....................................................... 160

16.2.7 Operations Timetable .......................................................................................... 161

16.2.8 Drilling Equipment and Productivity .................................................................... 162

16.2.9 Loading Equipment and Productivity ................................................................... 164

16.2.10 Ancillary and Support Equipment ........................................................................ 169

16.2.11 Total Fleet Required for the Mine ........................................................................ 170

16.2.12 Establishment of Equipment Lifetime .................................................................. 171

16.3 Description of Pit Operation and Infrastructure ............................................................. 171

16.3.1 Pit Drainage......................................................................................................... 171

16.3.2 Providing Electricity to Pit Operations ................................................................. 172

16.3.3 Storage and Preparation of Explosives ............................................................... 172

17 Recovery Methods ................................................................................................................... 173

17.1 Basic and General Criteria ............................................................................................ 173

17.1.1 Bases and Units Used ......................................................................................... 173

17.1.2 Definition of Capacity .......................................................................................... 173

17.1.3 Project Base ........................................................................................................ 173

17.1.4 Geography........................................................................................................... 173



17.1.5 Metallurgical Testwork......................................................................................... 174

17.2 Process Description ...................................................................................................... 175

17.2.1 Overview of Process ........................................................................................... 175

17.2.2 Crusher Circuit .................................................................................................... 176

17.2.3 Mill Feed Circuit................................................................................................... 178

17.2.4 Mill Circuit ............................................................................................................ 179

17.2.5 Gravity Circuit ...................................................................................................... 181

17.2.6 Pre-Oxidation, Leach and Adsorption Circuit ...................................................... 183

17.2.7 Carbon Elution Circuit ......................................................................................... 186

17.2.8 Tailings Disposal ................................................................................................. 187

17.2.9 Reagents ............................................................................................................. 188

17.2.10 Water Distribution ................................................................................................ 191

17.2.11 Electrical Power Supply....................................................................................... 192

17.2.12 Laboratory ........................................................................................................... 193

17.2.13 Plant Security ...................................................................................................... 193

17.2.14 Flowsheets and Layout ....................................................................................... 193

17.2.15 Main Process Equipments................................................................................... 197

17.2.16 Automation and Control....................................................................................... 197

18 Project Infrastructure ............................................................................................................... 203

18.1 General Criteria Adopted ............................................................................................... 203

18.2 Procurement and Distribution of Water ......................................................................... 205

18.2.1 Types of water ..................................................................................................... 205

18.2.2 Raw water ........................................................................................................... 205

18.2.3 Drinking Water..................................................................................................... 205

18.2.4 Process Water ..................................................................................................... 206

18.3 System and Distribution of Electricity ............................................................................ 206

18.3.1 Supply of Industrial Units to Mara Rosa .............................................................. 206

18.3.2 Distribution of Energy within Project Area ........................................................... 206

18.3.3 Fuel Supply System ............................................................................................ 207

18.4 Communication Systems (Internal and External) .......................................................... 207

18.5 Buildings - Maintenance Workshop, Office Buildings and Restaurant .......................... 207

18.5.1 Gatehouse ........................................................................................................... 207

18.5.2 Main Office Buildings........................................................................................... 207

18.5.3 Central Restaurant .............................................................................................. 208

18.5.4 Nurses Clinic ....................................................................................................... 208

18.5.5 Stores .................................................................................................................. 208

18.5.6 Central Maintenance Workshop .......................................................................... 208

18.5.7 Sanitary Waste and General Waste Disposal ..................................................... 208

18.6 Explosives Magazine ..................................................................................................... 209

18.7 Provisional Facilities (Implementation Period) .............................................................. 209

19 Market Studies and Contracts ................................................................................................ 210

19.1 Industry Trends and Pricing .......................................................................................... 210



19.2 Sale Strategy ................................................................................................................. 210

20 Environmental Studies, Permitting and Social or Community Impact ............................... 211

20.1 Physical Environment .................................................................................................... 211

20.2 Natural Environment ...................................................................................................... 211

20.2.1 Terrestrial Environment ....................................................................................... 211

20.2.2 Aquatic Environment ........................................................................................... 214

20.3 Social Environment ....................................................................................................... 217

20.4 Waste and Tailings Disposal, Site Monitoring and Water Management ....................... 219

20.4.1 Mining Area ......................................................................................................... 219

20.4.2 Tailings Basins .................................................................................................... 219

20.4.3 Water Usage ....................................................................................................... 219

20.4.4 Site Monitoring & Water Management ................................................................ 219

20.5 Permitting Requirements ............................................................................................... 220

20.6 Mine Closure ................................................................................................................. 222

21 Capital and Operating Costs ................................................................................................... 224

21.1 Mining Capital Cost ....................................................................................................... 224

21.1.1 Equipment Cost ................................................................................................... 224

21.1.2 Capital Cost for Mine Equipment......................................................................... 224

21.1.3 Pre-Production Services Cost ............................................................................. 227

21.1.4 Mine Services and Installations ........................................................................... 227

21.2 Plant Capital Cost .......................................................................................................... 229

21.2.1 Civil...................................................................................................................... 229

21.2.2 Mechanical Equipment ........................................................................................ 229

21.2.3 Platework............................................................................................................. 229

21.2.4 Metallic structures ............................................................................................... 229

21.2.5 Piping .................................................................................................................. 229

21.2.6 Electrical/Instrumentation Equipment .................................................................. 229

21.2.7 Cost of Installation (electrical and mechanical) ................................................... 230

21.2.8 Cost of the Water Supply .................................................................................... 230

21.2.9 Energy Supply ..................................................................................................... 230

21.2.10 Construction Management .................................................................................. 230

21.2.11 Miscellaneous Items ............................................................................................ 230

21.2.12 Main Equipment Costs ........................................................................................ 230

21.3 Infrastructure Capital Cost ............................................................................................. 235

21.4 Indirect Costs ................................................................................................................ 235

21.4.1 Studies and Construction Management .............................................................. 235

21.4.2 Miscellaneous...................................................................................................... 236

21.4.3 Sundry Items ....................................................................................................... 236

21.5 Sustaining Capital Costs ............................................................................................... 236

21.5.1 Mining Equipment................................................................................................ 236




21.5.3 Closure cost ........................................................................................................ 236

21.6 Schedule of Capital Costs ............................................................................................. 237

21.7 Mining Operating Cost ................................................................................................... 239

21.7.1 Overall Aspects ................................................................................................... 239

21.7.2 Basic Consumption and Cost .............................................................................. 239

21.7.3 Equipment – Hourly Costs................................................................................... 242

21.7.4 Operating Cost by Activity ................................................................................... 252

21.7.5 Consumption of Diesel Fuel ................................................................................ 253

21.7.6 Labour Cost and Requirement ............................................................................ 254

21.7.7 Operating Cost Summary .................................................................................... 258

21.8 Plant Operating Costs ................................................................................................... 260

21.8.1 Basis.................................................................................................................... 260

21.8.2 Overall Plant Operating Costs ............................................................................. 265

21.9 Summary of Operating Costs ........................................................................................ 266

22 Economic Analysis .................................................................................................................. 268

22.1 Cash Flow Assumptions ................................................................................................ 268

22.1.1 Mine Production Sequence ................................................................................. 268

22.1.2 Metallurgical Recovery ........................................................................................ 268

22.1.3 Metal Prices and Net Revenues .......................................................................... 268

22.1.4 Operating Costs .................................................................................................. 270

22.1.5 Capital Expenditures, Depreciation and Amortization ......................................... 270

22.1.6 Salvage Value ..................................................................................................... 271

22.1.7 Taxes................................................................................................................... 271

22.1.8 Working Capital ................................................................................................... 271

22.1.9 Closure Costs ...................................................................................................... 271

22.2 Financial Performance .................................................................................................. 272

22.2.1 Sensitivity Analysis .............................................................................................. 275

23 Adjacent Properties ................................................................................................................. 276

24 Other Relevant Data and Information .................................................................................... 277

25 Interpretation and Conclusions .............................................................................................. 278

25.1 Geology ......................................................................................................................... 278

25.2 Mining ............................................................................................................................ 278

25.3 Metallurgy and Mineral Processing ............................................................................... 279

25.4 Infrastructure ................................................................................................................. 279

25.5 Capital and Operating Cost Estimates .......................................................................... 280

25.6 Economic Analysis ........................................................................................................ 280

25.7 Risk ................................................................................................................................ 280

25.7.1 Introduction.......................................................................................................... 280

25.7.2 Risk Assessment ................................................................................................. 283

26 Recommendations ................................................................................................................... 285



26.1 Feasibility Study Work Program .................................................................................... 285

26.1.1 Geology ............................................................................................................... 285

26.1.2 Pit Geotechnics ................................................................................................... 285

26.1.3 Mining Study........................................................................................................ 285


26.1.5 Metallurgical Testwork......................................................................................... 287

26.1.6 Plant Design and Engineering ............................................................................. 287

26.1.7 Environmental and Community ........................................................................... 287

26.2 Feasibility Study Program and Budget .......................................................................... 287

27 References ................................................................................................................................ 289



List of Tables

Table 1.4_1 – Key Design Criteria for Mara Rosa Samples 5

Table 1.5_1 – Resource Summary 2011 Resource Estimate - 0.5 g/t cut off 6

Table 1.6_1 – Mineral Reserve Estimate (28 October 2011) 6

Table 1.6_2 – Input Parameters used for the Mineral Reserve Estimate (28 October 2011) 7

Table 1.7.2_1 – Production Scheduling 8

Table 1.7.3_1 – Staged Storage Capacity, Quantities and Implementation 9

Table 1.12_1 – Capital Cost Summary 17

Table 1.13_1 – Life-of-mine Economics (US$) 18

Table 4.2_1 – Mara Rosa Mining Concessions 31

Table 6.4_1 – Summary of Historic Drilling 42

Table 6.4_2 – Summary of Drillholes in the Posse database end 2009 42

Table 6.5_1 – Historic Grade Tonnage Estimates 43

Table 6.5.3_1 – Metallica Resource estimate (1.0 g/t Au cutoff) 44

Table 6.5.4_1 – Posse Deposit mineral resource estimate 45

Table 6.5.5_1 – HCS & AEFS Resource Estimate 2010 45

Table 6.6_1 – Summary of WMC Production at Mara Rosa, Posse Deposit Sulphides 46

Table 7.2_1 – Principal stratigraphic units of the Eastern Belt 50

Table 8_1 – Significant deposits in the Mara Rosa region 57

Table 10.5_1 – Historic SG measures 63

Table 12.2.1_1 – Certified values of standards 69

Table 12.3_1 – Coordinate conversion SAD69 to WGS84 70

Table 13.6_1 – Key Design Criteria for Comminution Testwork 80

Table 13.7_1 – Pre-Oxidation & Leach Design Criteria for Mara Rosa Samples 84

Table 13.8.4._1 – Acid Base Results Summary 89

Table 13.8.5.1_1 – Tails Multi Elemental Analysis – Main Composite 91

Table 13.8.5.1_2 – Tails Multi Elemental Analysis – Hanging Wall Composite 92

Table 13.10_1 – Design Criteria for Mara Rosa Samples 94

Table 14.2_1 – Summary statistics for raw assays in the mineralised zone 98

Table 14.2_2 – Summary statistics for 1m composite assays in the mineralised zone 99

Table 14.2_3 – Key modelling parameters 101

Table 14.2_4 – Modelling parameters common to all modelling runs 101



Table 14.2_5 – Modelling parameters common to all modelling runs 101

Table 14.2_6 – Key model statistics compared to the input data 102

Table 14.4_1 – Resource Summary 104

Table 15_1 – Mineral Reserve Estimate (28 October 2011) 106

Table 15_2 – Input Parameters used for the Mineral Reserve Estimate (28 October 2011) 106

Table 15_3 – Mineral Reserve Estimate Sources of Supporting Information (28 October 2011) 107

Table 15_4 – Break Even Grade Estimate 108

Table 16.1.5_1 – Hangingwall batter-berm configurations for different discontinuity dip angles 118

Table 16.1.5_2 – Footwall batter-berm configurations 119

Table 16.2.1_1 – Block Model Definition 121

Table 16.2.1_2 – Block Model Attributes 121

Table 16.2.1_3 – Grade and Tonnes Above Specified Cutoff 122

Table 16.2.1_4 – Mine Costs Distribution 122

Table 16.2.1_5 – Geotechnical Parameters Summary 122

Table 16.2.1_6 – Geometric and Economic Parameters for Pit Optimization 123

Table 16.2.1_7 – Mara Rosa Reserves - Results of Optimization in Whittle 124

Table 16.2.1_8 – Whittle Pit Tonnage and Grade Summary 126

Table 16.2.1_9 – Waste Summary - Pit US$1,100 / oz - Au < 0.50 g/t 127

Table 16.2.2_1 – Geotechnical Parameters of Pit Design 128

Table 16.2.2_2 – Tonnage and Grade Summary - Pit US$1,100 / oz - Au> 0.50 g/t 130

Table 16.2.2_3 – Waste and Resources Summary - Pit US$1,100 / oz - Au < 0.50 g/t 130

Table 16.2.3__1 – Operational Parameters for Mine Scheduling 131

Table 16.2.3_2 – Mine Scheduling 132

Table 16.2.4_1 – Volume of Stockpiled Materials 135

Table 16.2.4_2 – Waste Dump Dimensions 139

Table 16.2.4_3 – Program Implementation – Waste Rock Stockpiling 140

Table 16.2.4_4 – Program Implementation - Low Grade Ore Stockpiling 141

Table 16.2.4_5 – Program Implementation – Altered Rock Stockpiling 141

Table 16.2.5_1 – TSF Design Criteria 144

Table 16.2.5_2 – Storage Capacity and Raise Implementation 144

Table 16.2.5_3 – Embankment Statistics by Stage 147

Table 16.2.5_4 – Construction Quantities 150

Table 16.2.7_1 – Operations Timetable Criteria 161



Table 16.2.8_1 – Operations Timetable Criteria 162

Table 16.2.8_2 – Drilling Performance in Ore 163

Table 16.2.8_3 – Drilling Performance in Waste 163

Table 16.2.8_4 – Required Number of Drill Rigs 164

Table 16.2.9_1 – Productivity Data for Waste Rock 165

Table 16.2.9_2 – Productivity Data for Ore and Altered Rock 166

Table 16.2.9_3 – Transport Distances 167

Table 16.2.9_4 – Truck Velocities 167

Table 16.2.9_5 – Parameters Used in Calculating Truck Productivity 168

Table 16.2.9_6 – Transport Time and Quantity of Equipment Required for Ore Truck 169

Table 16.2.9_7 – Transport Time and Quantity of Equipment Required for Waste Truck 169

Table 16.2.10_1 – Number of Ancillary and Support Equipment Required 170

Table 16.2.11_1 – Main and Auxiliary Equipment Required 170

Table 16.2.11_2 – Necessary Support Equipment 171

Table 17.1.3_1 – Data Sources for Plant Design 173

Table 17.1.4_1 – Physiographic Data 173

Table 17.1.4_2 – Climate Data 174

Table 17.2.1_1 – Material Characteristics 176

Table 17.1.3_1 – Plant Crushing Criteria 177

Table 17.2.4_1 – Grinding and Classification Parameters 180

Table 17.2.5_1 – Gravity Circuit Parameters 182

Table 17.2.6_1 – Data Sources for Pre-oxidation, Leaching and Detoxification 184

Table 17.2.7_1 – Data Sources for Elution 187




Table 21.1.2_1 – Mine Equipment 225

Table 21.1.2_2 – Investment Schedule 226

Table 21.1.3_1 – Disbursement Schedule – Initial Capital Cost and Reinvestment 228

Table 21.2.12_1 – Plant Major Crushing Equipment List and Unit Costs 231

Table 21.2.12_2 – Plant Major Milling Equipment List and Unit Costs 232

Table 21.2.12_3 – Plant Major Gravity, CIP and Elution Sections Equipment List and Unit Costs 233

Table 21.2.12_4 – Plant Capital Costs 234



Table 21.3_1 – Infrastructure Capital Costs 235

Table 21.4.1_1 – Studies and Construction Management Costs 235

Table 21.4.2_1 – Miscellaneous Costs 236

Table 21.6_1 – Capital Cost Schedule 238

Table 21.7.2_1 – Main Equipment Hourly Fuel and Lubricant Consumption 240

Table 21.7.2_2 – Drilling Equipment – Costs and Life Cycle 240

Table 21.7.2_3 – Excavator Wear Parts - Life Cycle and Costs 241

Table 21.7.2_4 – Front Loader Wear Parts - Life Cycle and Costs 241

Table 21.7.2_5 – Wheel Tractor-Scraper and Tractor Wear Parts – Life Cycle and Costs 241

Table 21.7.3_1 – Hourly Cost Details for Rock Drill Rig (Ore) 243

Table 21.7.3_2 – Hourly Cost Details for Rock Drill Rig (Waste) 244

Table 21.7.3_3 – Hourly Cost Details for the Hydraulic Excavator (Ore) 246

Table 21.7.3_4 – Hourly Cost Details for the Hydraulic Excavator (Waste) 247

Table 21.7.3_5 – Hourly Cost Details for the Wheel Loader 248

Table 21.7.3_6 – Hourly Cost Details for the Haul Truck (Ore) 250

Table 21.7.3_7 – Hourly Cost Details for the Haul Truck (Waste) 251

Table 21.7.4_1 – Estimated Blasting and Fragmentation Unit Cost – 10m Bench 252



Table 21.7.5_1 – Total Diesel Fuel Consumption per Year 254

Table 21.7.6_1 – List of Operating Equipment and Labour 254

Table 21.7.6_2 – Mine Staff 255

Table 21.7.6_3 – Mine Maintenance and Warehouse Personnel 256

Table 21.7.6_4 – Total Mine Labour Cost 257

Table 21.7.7_1 – Summary of the Total Operating Cost for the Mine 259

Table 21.7.7_2 – Average Operating Cost 260

Table 21.8.1_1 – Plant Unit Costs 261

Table 21.8.1_2 – Plant Personnel Requirements and Cost Estimate 262

Table 21.8.1_3 – Staff Salaries (Admin and Plant) 264

Table 21.8.2_1 – Plant Operating Cost Estimate 265

Table 21.8.2_2 – G & A Cost Estimate 266

Table 22.1.3_1 – Average Historical Gold Prices 269

Table 22.1.3_2 – Adjustments to Gross Revenues 269



Table 22.1.4_1 – Average Unit Production Costs Full Production Years 270

Table 22.1.5_1 – Capital Cost Summary 270

Table 22.2_1 – Life-of-mine Economics (US$) 272

Table 22.2_2 – Life-of-mine Cash Flow 273

Table 22.2_1 – Cash Flow Sensitivity After Tax NPV at 5% Discount Rate, US$M 275

Table 25.7.1_1 – Qualitative Measures of Consequence 281

Table 25.7.1_2 – Qualitative Measures of Likelihood 282

Table 25.7.1_3 – Qualitative Risk Analysis Matrix 283

Table 25.7.2_1 – Summary of Risk Analysis 284



List of Figures

Figure 4.1_1 – Location of Amarillo’s Mara Rosa Properties 1

Figure 4.1_1 – Location of Amarillo’s Mara Rosa Properties 28

Figure 4.2_1 – Mining and Exploration Concessions over and around the Posse Deposit 30

Figure 5.4_1 – Mara Rosa, surrounding towns 34

Figure 7.1_1 – Summary geology of Brazil 48

Figure 7.1_2 – Mara Rosa Local Geology 49

Figure 7.2_1 – Mara Rosa District stratigraphic column 51

Figure 7.3_1 – Geology of the Posse Deposit 54

Figure 7.4_1 – Inclined longitudinal section of the Posse Deposit 56

Figure 10.1_1 – 2011 drill hole plan with the US$1,000 shell (brown) at 420RL 60

Figure 11_1 – ACME sample preparation flow chart 67

Figure 12.5_1 – MRP series holes 71

Figure 13.7.1_1 – Grind Size Recovery Relationship at pH12 after 24 hours 85

Figure 13.7.2_1 – Main Composite – pH Demand 86

Figure 13.7.2_2 – Hanging Wall Composite Lime Demand 86

Figure 14.2_1 – Posse Wireframes and drillholes 97

Figure 14.2_2 – Downhole and Directional semi-variogram, Au median indicator 100

Figure 14.2_3 – Comparison between 1m Composites and grades in the merged MIK model 102

Figure 14.3_1 – Comparison between 1m Composites and grades in the merged MIK model 103

Figure 14.4_1 – Grade Tonnage Curve 105

Figure 16.1.3_1 – Distribution of Field Strength Index Values for Mara Rosa Project 113

Figure 16.1.3_2 – Distribution of RQD Values for Mara Rosa Project 114

Figure 16.2.1_1 – Mara Rosa Resources - Results of Optimisation in Whittle 125

Figure 16.2.2_1 – Ramp Safety and Drainage 128

Figure 16.2.2_2 – Designed Pit with Single ramp system to both pits 129

Figure 16.2.3_1 – Mine Scheduling 133

Figure 16.2.3_2 – Pre-Stripping 134

Figure 16.2.3_3 – Period Year 7 - Final Pit Geometry 134

Figure 16.2.4_1 – Waste Rock Stockpile and Low-Grade Ore Stockpile Design Parameters 136

Figure 16.2.4_2 – Altered Waste Rock Stockpile Design Parameters 137



Figure 16.2.4_3 – Waste Rock Dump Layout 138

Figure 16.2.5_1 – TSF Options Study – Site Layout 143

Figure 16.2.5_2 – TSF Option 1 General Arrangement 146

Figure 16.2.5_3 – TSF Water Balance 155

Figure 16.2.5_4 – WSF Water Balance 155

Figure 17.2.14_1_1 – Plant Flowsheet Block Diagram 194

Figure 17.2.14_2 – Plant Process Flowsheet 195

Figure 17.2.14_3 – Plant Layout 196

Figure 16.2.4_3 – Master Plan – Site Layout 204

Figure 22.2_1 – Sensitivity Spider Diagram 275

Figure 25.7.1_1 – Framework for Risk Analysis 280


Mara Rosa Project, Goiás State, Brazil – 220810 Page: 1 Pre-Feasibility Study – 28 October 2011

1 SUMMARY

The Mara Rosa Project (the Project) is located in Goiás state, central Brazil. Amarillo Gold

Corporation (Amarillo), through its wholly owned subsidiary Metallica Brasil Ltda, currently

owns the Project and retained Coffey Consultoria e Serviços Ltda (Coffey Mining) to manage

and conduct a Pre-feasibility Study on the viability of mining the deposit from open pit mineral

resources and processing ore at an annual nominal production rate of 2.5 Mtpa to produce

gold doré. The work has been undertaken in collaboration with Amarillo and various other

independent consultants, including HCS & AEFS, BVP, Onix, Neotropica and Hidrovia.

Figure 4.1_1 Location of Amarillo’s Mara Rosa Properties



1.1 Property Description and Ownership

The Posse deposit is located in Goiás state, central Brazil, approximately 6 km north of the

town of Mara Rosa. It has a strike length of approximately 1,275 m, strike direction of 50° and

a dip of 40° to 45° to the northwest, terminates at surface and has a true thickness ranging

from 15 m and 30 m.

The property is centred at Latitude 13°58.395′ S, Longitude 49°10.690′ W (approximate

WGS84 coordinates 696880 mE, 8454530 mN, Zone 22 South). Presently, Amarillo holds a

property position totalling 80,833.9 ha of exploration leases and 2,552.62 ha of mining leases

as part of the Mara Rosa Project

The region is characterized by tropical savannah of low to moderate topographic relief ranging

from approximately 400 m to 500 m above sea level. Much of the area has been cleared for

farming and as a result is open savannah grassland.

1.2 Geology and Mineralization

The Mara Rosa District is situated within the Goiás Magmatic Arc (“GMA”) which forms part of

the Tocantins physiographic province, an intercratonic mobile belt that separates the

Amazonas and São Francisco cratons, located to the northwest and southeast respectively.

The GMA is a 100 km wide, northeast-trending granite-greenstone terrane that extends

approximately 700 km. The geology in the Mara Rosa District is principally delineated by three

northeast-striking, moderately to steeply northwest-dipping belts of metamorphosed volcano-

sedimentary and associated intrusive rocks. These belts, referred to as the Western, Central,

and Eastern Belts, are separated by broad zones of tonalitic orthogneiss.

The Eastern Belt is bounded to the southeast by the Rio dos Bois fault, which also defines the

south eastern limit of the GMA.

Amarillo’s land position within the Mara Rosa District primarily covers the Eastern Belt

greenstone assemblage. The Eastern Belt, which in general strikes to the northeast and dips

moderately to steeply to the northwest, has a maximum thickness of 6 km. Surface

topography over the belt is characterized by moderate relief and locally dissected drainages

that follow lithologic or structural weaknesses. Depth to fresh bedrock is generally shallow,

ranging from 0 m to 15 m. The upper portion of the weathered profile consists of clay-rich

latosol and saprolite derived from the underlying bedrock.

Several significant mineral deposits occur within 50 km of Mara Rosa including the Posse

gold deposit, the Zacarias gold-silver deposit and the Chapada copper-gold deposit, in

addition to numerous historic prospects and garimpos.

Alteration and mineralization at Posse are characterized by silicification, sericitization, K-

feldspar flooding and pyritization. Carbonatization, usually as ankerite, is present, though

relatively minor. Pyrite is the dominant sulphide occurring as 1% to 5% finely disseminated

grains. Accessory metallic minerals (which typically compose less than 1% of the rock)



include calaverite (a refractory gold telluride), frohbergite (iron telluride), chalcopyrite,

molybdenite and sphalerite. Gold occurs as 10 to 100 µm-sized native grains along the

margins of silicates and in association with pyrite and frohbergite. In general, gold grade tends

to show an overall positive correlation with intensity of silicification and total sulphide content.

1.3 Status of Exploration

Historically Western Mining Corp. (“WMC”) operated a small open pit mine at the project site

during the 1990s. Two pits, Posse South and Posse North, were developed over a 5 year

period and the ore was processed on-site. As of November, 2006, the mine and mill site had

been reclaimed and no site infrastructure remained. No significant environmental liabilities are

known to exist at the former mine site and it is understood that the required remediation for

mine closure had been met and accepted by the appropriate government agencies

A number of drilling campaigns have been completed on the Property by BHP Billiton from

1982 until 1987, by WMC from 1988 until 1995, and by Metallica in 2002. In addition Amarillo

has completed three drilling programs, one in 2005/2006 and another in 2008 and the third

from October 2010 to March 2011. In all, the drillhole data base contains 277 drill holes for a

total of 33,600.0 m.

1.4 Metallurgical Testwork

Metallurgical testwork was undertaken on representative samples from the Mara Rosa Project

as part of the pre-feasibility study (PFS).

The Mara Rosa metallurgical samples comprised of a free milling component of gold (~75%)

and a refractory component, which may be associated with sulphides as well as tellurides

(~25%).

Gold recoveries in excess of 93% were readily achieved using conventional carbon in leach

(CIL) technology with the addition of a simple pre-oxidation stage (3 agitated tanks) prior to

the CIL circuit to account for the refractory gold telluride component in the mineralisation.

Cyanide consumption was low at 0.26 kg/t, whilst lime requirements were slightly higher than

average at 1.93 kg/t. The slightly higher lime requirement is needed to raise the pH to 12 to

accelerate the gold telluride oxidation process.

A review of the previous metallurgical testwork found many of the results to be inconsistent

and less successful due to a lack of understanding of gold telluride chemistry, which although

not overly complex, is rarely seen in practice with less than a handful of plants treating such

materials worldwide.

Mineralisation containing gold tellurides simply needs to have an allowance for the oxidation

of the gold telluride, using agitated tanks similar in size to the CIL tanks. The Mara Rosa

mineralisation needs to be ground to a P80 size of 45 µm prior to a pre-oxidation stage of 12

hours and a CIL stage of 24 hours. The overall plant residence time of 36 hours is strongly



influenced by the relatively slow process of gold telluride oxidation. Given recent

improvements in the understanding of aqueous sulphide and telluride oxidation, it is

considered that there remains considerable scope for a reduction in the residence time of the

plant with the completion of testwork to feasibility level.

There were three mineralogical domains identified; main, hanging wall and foot wall, although

there were no significant differences between the domains from a metallurgical perspective.

The samples tested were in the soft to medium range of competency in terms of milling and

displayed no viscosity issues even at finer grind sizes. The abrasion index of the samples

was in the medium to high range.

It is proposed that the process flowsheet would include primary crushing followed by

secondary and tertiary crushing in closed circuit. Tertiary crushed material would then feed a

primary mill and secondary mill utilising cyclone classification to achieve a P80 grind of

approximately 45 µm.

This material would be pre-oxidised at a pH of 12 for a period of 12 hours in agitated tanks

and then leached under conventional CIL conditions for 24 hours during which an average of

93% of the gold would be dissolved and adsorbed onto activated carbon. The adsorbed gold

would then be eluted using a conventional desorption plant.

Tailings would be thickened to recover a significant proportion of the cyanide in the process

solution, with the remaining thickened pulp detoxified to remove residual free cyanide prior to

deposition in a tailings storage facility.

Table 1.4_1 shows a summary of the key design criteria based on the metallurgical testwork

results for the Mara Rosa samples.



Table 1.4_1

Mara Rosa Project

Key Design Criteria for Mara Rosa Samples

Criteria Unit

Annual throughput tpa 2,500,000

Availability % 90.0%

Instantaneous throughput tph 317

Bond rod mill work index kWh/t 13.4

Bond ball mill work index kWh/t 13.0

Abrasion index 0.3426

P80 grind size µm 45

pH set point 12.0

Pre-oxidation time h 12

Cyanidation leach time h 24

Cyanide consumption kg/t 0.26

Lime consumption kg/t 1.93

Thickener settling rate t/m2/h 0.50

Gold head grade g/t Au 1.47

Gold in residue (design) g/t Au 0.10

Gold recovery (design) % 93.2

Gold in residue (optimum) g/t Au 0.06

Gold recovery (optimum) % 95.9

1.5 Mineral Resource Estimates

Since Amarillo has controlled the project several Mineral Resource Estimates have been

made. The updated Independent Mineral Resource Estimate presented herein was published

on 30th July 2011 and is based upon:

previously compiled historic data

New drill data

Updated topography

The use of three mineralised domains: HW (Hanging Wall), MAIN (MAIN) and FW

(Footwall) through the use of three sets of wire frames.

a revised estimation methodology; and,

Re-classification of the Estimate.

The Mineral Resource Estimate, summarized below in Table 1.5_1, was estimated utilizing

the Median Indicator Kriging method, without top cutting of gold grades. Blocks were

classified as Measured, Indicated or Inferred based on data density defined via a long section

in the plane of the vein (i.e. pierce point density).



Table 1.5_1

Mara Rosa Project

Resource Summary 2011 Resource Estimate - 0.5 g/t cut off

Author Category Tonnes Au grade g/t Ounces

AEFS/HCS 2011 Measured 5,462,000 2.04 358,500

AEFS/HCS 2011 Indicated 15,393,000 1.65 815,200

AEFS/HCS 2011 Inferred 3,629,000 1.34 156,400

1.6 Mineral Reserve Estimates

Table 1.6_1 shows the Mineral Reserve estimate, based on a Mineral Resource cutoff grade

of 0.5 g/t Au. The Mineral Reserve is included within the declared Measured and Indicated

Mineral Resource and is declared inclusive of approximately 0.5 Mt of dilution at an average

grade of <0.2 g/t.

Table 1.6_1

Mara Rosa Project

Mineral Reserve Estimate (28 October 2011)

Classification Tonnes (Mt) Au grade (g/t) Contained Gold (Moz)

Proven Mineral Reserve 5,366,400 1.97 339,600

Probable Mineral Reserve 11,750,400 1.60 606,600

Total Mineral Reserve 17,116,800 1.72 945,200 The tonnes and grade reported here is Run of Mine. Application of the plant recovery factor reduces the recoverable

gold to 869,600 oz.

Rounding has been applied.

The Mineral Reserve estimate has been determined and reported in accordance with the CIM

Definition Standards (2010).

The reported Mineral Reserve has been compiled under the supervision of João Augusto

Hilário, MAIG, an employee of Coffey Consultoria e Serviços Ltda.

A summary of the main input factors used in estimating the Mineral Reserve are shown in

Table 1.6_2.



Table 1.6_2

Mara Rosa Project

Input Parameters used for the Mineral Reserve Estimate (28 October 2011)

Description Units Value

Gold price US$/oz 1,100

Mineral resource Au cut off grade g/t 0.5

Mining method Open pit

Annual production rate Mtpa 2.5

Mining operating cost US$/t ore 12.59*

Processing operating cost US$/t ore 9.73 **

G&A operating cost US$/t ore 1.83

Mining dilution % 3

Mining recovery/loss % 97

Plant recovery % 92

Project capital cost US$M 181

Sustaining capital cost US$M 19

Royalty % 2

Pit slope degrees 55° HW 40° FW

Strip ratio 8:1 * Mining operating costs are quoted in this table inclusive of the Year 0 pre-stripping.

** Processing operating costs are estimated for a plant design throughput of 2.5 Mtpa.

1.7 Mining Methods

1.7.1 Geotechnical

The geotechnical work to date, and data provided, have been reviewed and summarised.

Coffey Mining’s review of these works suggests that they are sufficient for PFS level.

Preliminary pit slope design parameters have been derived using the Haines-Terbrugge

empirical approach for the hangingwall; the footwall slope angle and design parameters are

considered to be likely to be controlled by the dip of the orebody. In this study a conservative

angle less than the dip of the orebody has been applied, with anticipated improvement once

further drilling and testwork has been completed. The hangingwall has also been assessed

for potential toppling failure; with the currently available data the modelled factors of safety for

this wall are greater than 1.3.

Groundwater and seismic effects, and operational considerations such as blasting, may affect

the PFS geotechnical slope design.

Recommendations for drilling, test work and geotechnical studies to advance the geotechnical

aspect of the project to Feasibility Study level are given in Beer (2011), together with a cost

estimate for the geotechnical test work and geotechnical studies recommended.

1.7.2 Mining Study

The Mara Rosa deposit will be mined in a conventional open pit using 45 t and 90 t haul

trucks for ore and waste respectively. Proven and Probable Mineral reserves have been



estimated by first modelling an optimised pit using the parameters listed in Table 1.6_2. Mine

design and scheduling has applied the further constraints imposed by minimum working

areas, ramp widths, curvature and gradient.

The production schedule based on the mineral reserves will produce up to 2.5 Mtpa over a

seven year life of mine with pre-stripping and production ramp up in years 0 and 1. Table

1.7.2_1 shows the proposed production schedule.

Table 1.7.2_1

Mara Rosa Project

Production Scheduling

Unit 2013 2014 2015 2016 2017 2018 2019 2020 Total

Mine production

kt 200 2,298 2,404 2,451 2,440 2,517 2,444 2,361 17,117

Contained gold

oz 6,500 119,400 145,500 145,700 133,000 124,500 129,100 141,600 945,200

Grade g/t 1.01 1.62 1.88 1.85 1.69 1.54 1.64 1.87 1.72

1.7.3 TSF

The design for the Tailings Storage Facility (TSF) and related Water Storage Facility (WSF)

for the Mara Rosa Gold Project has been aimed at optimising tailings storage capacity by

maximising tailings density; and reducing environmental and societal impact.

Based on the site selection study that was carried out in June 2011, Site 1 has been used as

the preferred area in which to develop the prefeasibility study design. To the east of the TSF

basin, there is a potential water storage dam site to provide storage capacity for the decant

water and stormwater collected from the TSF.

The information that has been provided to Coffey Mining, apart from the topographical

mapping, included the conceptual master plan (August 2011), the hydrogeological report,

process mass balance (September 2011) and a preliminary water balance (20 September

2011). The TSF design is to be based on a slurry with 59% solids (W/W) that is expected to be

deposited at 1.25 t/m³ with a beach slope of 1%. Mine production is expected to be 20 Mt

over a period of eight years. The design was carried out in accordance with ANCOLD

guidelines.

Construction of the TSF will be in four stages over the life of the mine, and will take advantage

of the progressive development of the open pit with the quantities of deposited tailings and

construction material as listed in Table 1.7.3_1:



Table 1.7.3_1

Staged Storage Capacity, Quantities and Implementation

Stage No. Dry Mass Stored

(t) Crest

RL (m) Embankment Fill*

Volume (m³) Implementation

(Years after start)

1 4,073,000 465.0 787,000 Start up

2 3,877,000 470.5 560,000 2.04

3 3,969,000 474.5 463,000 4.00

4 4,274,000 478.0 570,000 6.00 *Includes volumes for decant causeway.

Extensive volumes of waste rock are intended to be used in the construction of the

downstream portion of the perimeter embankment with an upstream core of clay material

separated by a geofabric and coarse stone filters. The tailings slurry will be discharged from

the crest of the perimeter embankment through a series of spigots off a delivery pipeline.

Lengths of rockfill causeway will lead from the perimeter embankment to the three decant

towers and the power supply cables and pipelines for the decant pumps will run on the

embankment crests, which will also provide vehicle access. Water management for the TSF

will be carried out using the decant pumps in the towers to recover both supernatant and

seepage water from the tailings and conveying this to the WSF to the east of the TSF, from

where this water can be returned to the plant for reuse.

The high flood level of the WSF is constrained at relative level (RL) of 460m by the Stage 4

TSF embankment, with the full supply level (FSL) at RL 459 m. A saddle on the right bank

(north eastern side) of the valley is where the spillway will be situated. The WSF provides

storage capacity of up to 514,000m³ with a maximum surface area of 12.3ha. The zoned

embankment 12m high will have a central clay core and gravelly clay upstream and

downstream faces, and will also incorporate a downstream rock fill toe to provide drainage

and stability. The eastern saddle embankment will have a crest width adequate for the site

access road and incorporate box culverts to form the spillway to this storage. Built under the

main embankment will be an outlet pipe, controlled at the downstream end by valves that will

regulate the flow to the plant.

A thickened slurry of tailings with 59% solids will be discharged sub-aerially onto a beach from

a number of spigots located along the main embankment crest. Management of the TSF

requires the implementation of recommended action dependent on the monitoring results.

Whilst most of the monitoring will be visual, less frequent exercises involve:

measurement of standing water levels in bores;

levelling of embankment settlement monuments; and

laboratory testing of ground water samples from bores.

Although the design of the TSF and WSF mitigate many environmental risks, there are

residual hazards that occur at specific times during the life of the TSF and these are:



During construction:

dust from haul and inspection vehicles;

plant exhaust fumes and spill of engine oil;

noise generated by the trucks and construction plant; and

unnecessary removal of vegetation.

During operation:

elevated water table;

dust from the inspection vehicles;

windblown silt and sand from the dry TSF beach;

uncontrolled release of tailings, and

spill from the return water pipeline.

Post mining:

windblown dust generation before rehabilitation;

decreasing pH of ground water; and

seepage of poor quality water from the TSF.

These have been identified with a view to implementing measures to mitigate the impacts

during project implementation, if required.

Additional investigations are recommended and include:

Geotechnical investigation for the TSF and WSF.

Drilling and equipping of the monitoring bores.

Drilling and permeability testing of investigation bores.

Laboratory testing of tailings and waste rock geochemistry and geotechnical properties.

Confirmation of geotechnical parameters for the embankment construction materials.

The next stage of design work is the feasibility design report for the TSF and WSF which will

provide a comprehensive document that will assist in progressing the project.

1.8 Recovery Methods

During plant operations from 1992 to 1995, Western Mining Corporation were aware that

recoveries were declining as the pit was deepening and less oxide ore and more sulphide ore

(with tellurides) was being processed. The recovery in the final month of production before

closure due to the low gold price was only 83%.



A large number of testwork campaigns have been carried out over the years, mostly

misguided as the proponents did not have the necessary experience to design programmes to

allow for the presence of the tellurides. High extractions were only achieved (Western Mining

Laboratories and Testwork Technologies) when oxidation was applied in the form of high

additions of calcium hypochlorite, implying a very expensive process route.

Under the guidance of metallurgical consultants with personal experience of deposits

containing tellurides (in particular, the Finiston mine in Kalgoorlie operated by Kalgoorlie

Consolidated Gold Mines), a programme involving oxidation on a more practical level in terms

of costs was formulated.

As a result the laboratory testwork indicated that recoveries of 93% could be obtained from

the two main ore types (Main and Hanging Wall). It is considered that this applies to more

than 97% of the gold content of the deposit. The ore type FW, of lower grade, with less than

3% of proven or probable reserves gave recoveries in the order of 86% under the same

conditions.

The final process route is again applying strong oxidation but by milling to a P80 of 45 μm and

exposing the ground mineral to oxidation of the pulp with the injection of low grade oxygen

gas (delivered by an inexpensive PSA plant) for 12 hours at a pH of 12, an economic process

route with a high recovery in the subsequent cyanidation stage has been achieved.

The all mill work index is 13 kWh/t (considered reasonable) and therefore the fine grinding

does not involve excessive costs. The PSA oxygen plant will also consume some additional

energy (325 kW) and a calculation of the overall extra energy cost per tonne of ore results in a

value of 7 kWh/t processed. This includes the extra total agitation time of 36 hours against the

more usual 24 hours. Cyanide costs will be modest at a consumption rate of around 0.26 kg/t.

Lime consumption at 1.93 kg/t can be considered high but this is a relatively cheap chemical

and readily obtained in the area.

Thus the final plant will include a grinding stage composed of a 4 MW primary mill in open

circuit followed by secondary mill, also 4 MW, in closed circuit with cyclones to produce a final

milled product with a P80 of 45 μm. Lime will be added in solid form with the feed to the

primary mill to maintain a pH of 12 as measured in the leach circuit. This pulp will be

contacted in 3 equal sized agitated vessels (2070 cubic metre capacity) where a pump will

circulate pulp through a contactor where oxygen from a PSA unit will be added to obtain

dissolved oxygen contents in the order of 20 ppm. The total average residence time of the

pulp in this oxidation stage will be 12 hours. The pulp overflowing from the third tank will enter

the leach/CIL circuit where cyanide and activated carbon is present to dissolve and adsorb

the gold in a traditional six stage system each with a 4 hour residence time making a total

cyanidation residence time of 24 hours. The carbon will be removed at intervals and the gold

extracted using a typical Zadra type elution circuit. The reject pulp will be thickened for

solution recirculation before being subjected to a detoxification stage before being pumped to

the tailings pond.



1.9 Project Infrastructure

Power will be supplied by a 65 km 132 KV power supply line and substation in the nearby city

of Porangatu.

Water requirements will be largely supplied by return water from the tailings dam and from

dewatering the pit with make-up quantities coming from wells (for good quality water used to

supply the potable water plant as well as for the make-up of reagents and pump seal water)

as well as water from the local river during the wet season.

A preferred tailings storage facility site has been located south west of and within 1 km of the

plant site. A wet slurry tailings storage facility has been designed for 20 Mt of tailings to be

constructed in four stages over the life of mine followed by vegetation coverage for final TSF

closure.

Lime will be supplied locally, cyanide in solid form from the Candeias plant in the state of

Bahia whilst mill balls will be imported from Chile.

Other infrastructure including an operations camp, surface workshops and warehouse,

canteen and administration buildings have been developed to support cost estimation and

development of general arrangements for the project.

1.10 Market Studies and Contracts

The Mara Rosa Project will produce gold bars containing about 95% gold for sale after

refining to financial institutions in Brazil or internationally.

The gold price in 2011, achieved a record spot trading value of more than US$1,800/oz.

The gold price that was utilized for the base case cash flow analysis was US$1,200/oz for the

life of the project, which approximates the prevailing three year trailing average for gold at the

time of the study.

1.11 Environmental Studies, Permitting, and Social and Community Impact

1.11.1 Physical and Natural Environment

The climate of the Mara Rosa region is characterised by a hot, wet summer and a cool, very

dry winter. Average temperatures in May/June are a pleasant 24 °C whereas average

temperatures in August/September reach 28 °C. Average rainfall as measured at Estrela de

Norte (30 km to the north) is 1,679 mm (period 1971 to 2010). Maximum humidity over 80%

is reached in April and December, while minimum humidity of about 40% is experienced in

July and August.

Studies of the natural environment were undertaken to establish baseline conditions for flora,

fauna and the aquatic environment at the Mara Rosa site for use in identifying possible

impacts from planned gold mining operations.



The project is situated within the Cerrado Biome, an immense tropical savannah that

constitutes Brazil’s second largest plant formation found in the Central Plateau regions

covering some 23% of the country. This biome is one of a select few on earth displaying both

high biodiversity and threatened ecological status. The high human population of the area

has resulted from socioeconomic intervention intended to transform the region into the

country’s breadbasket.

Satellite imagery has indicated four distinct vegetative communities within the project area

and biological studies have been undertaken for each community.

The vegetation survey has evaluated any plant formations of special interest, and their

preservation. The ecological functions and the environmental services of the vegetation were

studied to ascertain any species protected by law. The aim of the study has been to develop

a baseline database of information for the mining company to plan a sustainable method of

mining, and help develop mitigation measures and/or create conservation areas where

preservation or recovery will be implemented. This process must be undertaken in

accordance with Ordinances 14/2001 and 15/2001.

The studied fauna of the area comprise some 16 mammal species, 22 reptile and amphibian

species, and 79 bird species. Of the mammal species, two are considered vulnerable to

extinction and will require further study. Although none of the reptile or amphibian species are

vulnerable, the inventory data may be too few and inadequate for population estimation and

conservation status. The bird species are highly dependent on the forest environments and

will require careful monitoring of population levels so that rapid environmental changes

caused by human activity do not adversely affect avifauna communities.

The aquatic environment includes three sub-basins, namely the Upstream Basin for the Ouro

River; the sub-basin for Lambari Creek and the sub-basin for the Antas River.

Surface water quality in the old mine pits is satisfactory and can be safely discharged. Study

of the phytoplankton and zooplankton communities showed a relatively high taxonomic

complexity and a typical tropical freshwater state for retained water.

The benthic macro-invertebrate community was found to be extremely poor and as such

draining the pits will not cause problems of general diversity.

Heavy metal concentrations were identified in sediments from the pits and it is recommended

that sediment at the bottom of these pits should be transferred and impounded in the rejects

dam, thus reducing the possibility of contaminating local surface water courses.

1.11.2 Hydrology and Hydrogeology

Mining was last carried at Mara Rosa in 1995; significant environmental changes from the

natural state include partially re-vegetated waste and tailings heaps and deforestation but



natural regrowth has hidden the worst of the degradation. The two old mining pits are now

water-filled.

Using the updated Master Plan (site layout) a plan for the monitoring and sampling of surface

and groundwater has been developed that will allow interpretation of water flows and quality

for compliance with legal requirements and permitting.

Mara Rosa town, 5 km from the mine site, is supplied with water from a strong aquifer but the

bedrock geology in the planned mine site is not favourable for adequate groundwater

supplies.

The paucity of seasonal hydrological data for the study area has precluded detailed estimates

of available surface water in the Rio do Ouro basin. However, preliminary estimates suggest

that even after required minimum flow and allowances for agricultural purposes, some

626 m3/h may be available; this is far in excess of the projected project needs of 114 m3/h.

Previous surface and groundwater chemical analyses have not detected any contaminant

values in excess of CONAMA Class II purity limits.

Recommendations are made to protect existing monitoring wells from airborne contamination.

The old tailings from a different ore body (Zacarias) in the Baribras area may present an acid

rock drainage risk which should be investigated and a mitigation plan implemented.

Continuing monitoring will be required within the mine area for diagnosis, assessment and

mitigation plans for environmental damage especially adjacent to the APP areas bordering

streams and rivers. The information will be integrated within the Environmental Impact Study.

1.11.3 Social Environment

A preliminary socio-economic study of the demographics, land use, production and

economics and quality of life has been undertaken using standard indicators and both direct

and indirect survey methods. The economic base of Mara Rosa is small scale agriculture and

livestock. The area is well serviced for a broad range of normal social and community

facilities and is considered to have a high human development index when compared to other

areas of Brazil.

Future mining operations are viewed favourably due to the generation of employment and

income. However these operations will require necessary improvements to the physical and

social infrastructure. Licensing will be required. The environmental licensing will require a

more detailed archaeological survey and report.

1.11.4 Waste and Tailings Disposal, Monitoring and Water Management

The waste water from the initial and continuing pit dewatering will be pumped for storage in

the WSF prior to use in the processing plant.



The tailings basins will be constructed with impermeable floors in order to contain and retain

any chemical compounds used in the process plant, thus facilitating neutralisation.

Daily water consumption of 5,000 m3 will be largely supplied from pit dewatering and tailings

dam return water. Supplementary water is expected to be largely surface water from streams

for which a more detailed hydrological study will be required.

Air and water quality will be indicators to monitor during the project life. The objective to

restrict human intervention outside the service area must also be monitored through indicative

studies.

1.11.5 Permitting

There are numerous laws and regulations governing mining and environmental issues at the

federal, state and municipal levels of authority in Brazil. Amarillo has engaged a local

Brazilian Environmental company to undertake all environmental permitting requirements for

the project. It is Coffey’s understanding that currently there are no outstanding permits for the

project

1.11.6 Mine Closure

A fundamental part of the permitting process is a Mine Closure Plan. A general view of what

will be required, and is in the process of being formulated for permitting purposes, includes:

Dismantle and sell or remove equipment for scrap, bury foundations, cover with soil and

revegetate;

TSF closure through capping, soil cover and revegetating;

Monitoring of water reservoir with detoxification if required until chemical composition is

acceptable for agricultural use (possibly two years);

Contouring, covering and revegetating waste rock dumps; and

Passive flooding of the mine pits with perimeter fencing and water quality monitoring.

1.12 Capital and Operating Costs

Coffey Mining has compiled a capital cost estimate with a precision of ±25% for the Pre-

Feasibility Study. The capital cost estimate is based on:

Major equipment costs from supplier quotes;

Contractor quotes for pre-stripping costs;

Unit costs from contractor quotes for TSF phased construction;

Power utility quote for powerline and sub-station construction with rebate against future

power operating bills;



Unit costs for mine infrastructure elements including workshops, warehouse and office.

Capital costs include direct and indirect costs for the mine, process plant and infrastructure.

Project capital costs total US$189.5 M. The total indirect cost is US$31.6 M and includes

studies and management, rebate from power utility, insurance, contingency and initial working

capital. A 10% contingency is placed on initial direct and indirect capital costs for the mine,

plant and surface infrastructure. The total contingency allowance for the project is US$16.2 M.



Table 1.12_1

Mara Rosa Project

Capital Cost Summary

Capital Cost Item Initial US$ M Sustaining

US$ M Total US$ M

Direct

Mine Development (pre-strip) 14.2 14.2

Mining Equipment 24.6 9.7 34.3

Processing (inc. TSF) 91.4 9.6 101.3

Infrastructure 13.9 12.6

Direct Subtotal 144.1 19.3 163.4

Indirect

Studies and Management 14.0 14.5

Miscellaneous 3.4 3.4

Rebate from CLEG (7.9) (7.9)

Insurance 0.4 0.4

Process Contingency 16.2 16.2

Initial Working Capital 5.5 5.5

Indirect subtotal 39.5 (7.9) 31.6

Total Capital (excluding working capital) 178.1 11.4 189.5

Total Capital 183.6 11.4 195.0 Rounding has been applied.

Coffey Mining, Amarillo and Onix have compiled operating costs from:

Production schedule tonnes;

Equipment operating hours and unit costs;

Personnel requirements; and Brazil unit costs for materials, consumables, services and

labour.

All costs have been estimated assuming Amarillo will be the operator and using an exchange

rate of R$1.9 = US$1.0. Blasting operations are assumed to be conducted by a contractor.

The Project operating costs include fixed and variable costs for mine production, plant

production, tailings management and general and administrative services for the operation.

Mine operating costs are estimated at US$12.59/t of ore, with a strip ratio of 8:1. However the

cost of waste rock pre-stripping in Year 0 has been capitalised for the purposes of economic

analysis. Consequently, mine operating costs average US$11.85 /t of ore during the

production Years 1 to 7 (excluding pre-stripping in Year 0 that are capitalised), with an

operational strip ratio of 7.4:1.

Plant operating costs, at a design processing rate of 2.5 Mtpa, total US$9.73 /t ore processed

including tailings disposal, and G&A costs average US$1.83 /t ore. These costs, pro-rated for

varying annual mine production rates and applied in the financial analysis, are US$9.78 /t

G&A costs and US$1.90 /t of ore processed respectively.



Total cash operating costs are US$21.63 /t ore or US$464 /oz of gold produced, averaged

over life of mine Years 1 to 7.

1.13 Economic Analysis

The overall economics of the Mara Rosa Project have been evaluated using conventional

discounted cash flow techniques based on the production schedules, capital expenditures and

operating costs. The following key parameters were integral to the cash flow model and the

economic results:

The base case metal price was $1,200/oz gold;

The analysis is based on 100% equity financing with no debt component;

All costs and revenues are reported in “real” or constant US dollars without escalation;

An income tax rate of 25% was applied, based on the general understanding of Brazilian

income tax laws;

Provision was also made for the Brazilian Social Contribution Tax of 9%.

A cash flow model incorporating Project and life of mine production, capital costs and

operating costs indicates that the Project has an after tax NPV of US$178.5 M at a discount

rate of 5%. A summary of the life of mine economics is presented in Table 1.13_1.

Table 1.13_1

Mara Rosa Project

Life-of-mine Economics (US$)

Tonnes of Ore Processed (000s) 17,117

Average ROM Grade, g/t Au 1.72

Gold Ounces Sold (000s) 869,592

Total Revenues (M) 1,044

Revenue per tonne 61.29

Mining Cost per tonne (Year 1 to 7) * 11.85

Processing Cost per tonne (at design 2.5 Mtpa) 9.73

G&A Cost per tonne (at design 2.5 Mtpa) 1.83

Processing Cost per tonne (at scheduled plant throughput) 9.78

G&A Cost per tonne (at scheduled plant throughput) 1.90

Operating Cost per ounce 464

OperatingCost per ounce (including refining and Royalties) 524

Capital Costs (millions) 189.5

Initial Working Capital (millions) 5.5

Net Present Value at 5% (pre tax, M) 283.1


Net Present Value at 5% (after tax, M) 178.5


Internal Rate of Return (after tax) 26.6%

Payback Period (after tax, production years) 3.0 * Note that mining cost, including pre-stripping in Year 0, is US$12.59 per tonne



A sensitivity analysis considering positive and negative variations of up to 30% were applied

independently to: gold price, operating cost and capital cost. The results of the sensitivity

analysis demonstrate that the project is most sensitive to variation in gold price. Initial capital

cost had the least impact on the sensitivity of the NPV.

While the base case gold price for the financial analysis was US$1,200/oz, the 30% positive

variance on gold price represents US$1,560/oz and at this metal price the NPV is

US$332.2 M.

1.14 Conclusions and Recommendations

The results of the Pre-Feasibility Study indicate that the Mara Rosa Project, under the

assumptions in this study, is a viable open pit mining and mineral processing operation.

A risk assessment based on consequences and likelihood has concluded there are no

extreme risks. High risk events include:

Mining costs under-estimated.

Pit slope incorrect.

Reserve grade over-estimated.

Mitigation measures involve further study and more detailed cost estimation during the

feasibility study and careful grade control measures during mining.

Similar mitigation is proposed for other lower risk events.

There is opportunity to increase mineable resources through exploration drilling into the

Inferred Mineral resources that separate the North and South Pits in the mine design.

Conversion of these resources into higher confidence categories would improve mine life as

well as reduce the strip ratio and thus operating costs.

Further opportunity may be realised if the gold price continues at a level above the base case

price adopted for this study.

A Feasibility Study is recommended for the Mara Rosa Project.

The recommended work plan for the Feasibility Study has already begun with a geotechnical

drilling program and further metallurgical testwork.

The following activities are recommended:

Drilling (US$2.4M) to collect data and samples for:

resource model update

geotechnical characterization and condemnation of tailings and plant site locations



samples for metallurgy and tailings test work

Metallurgical testwork program (US$0.4M).

Geotechnical testwork program(US$0.4M)

Hydrogeological study (US$0.4M)

Tailings site testwork program (US$0.4M)

An updated Mineral Resource Model incorporating exploration data to improve

confidence in Mineral Resources (US$0.05M).

An updated mine design and mine schedule incorporating new hydrogeological, and

geotechnical data testwork. (US$0.5M).

Feasibility study including process and infrastructure design, engineering, capital and

operating cost estimation and financial analysis incorporating results of the geotechnical,

hydrogeological, mine design and mine schedule and metallurgical test work (US$1.5M)

Field expenses to continue with the environmental base line study, property

maintenance, field staff and overheads (US$1.0M)

The recommended feasibility work plan will require a budget of approximately US$6.3M.



2 INTRODUCTION

The Mara Rosa Project (the Project) is located in Goiás state, central Brazil. Amarillo Gold

Corporation (Amarillo), through its wholly owned subsidiary Metallica Brasil Ltda, currently

owns the Project, which is at the Development stage, and retained Coffey Consultoria e

Serviços Ltda (Coffey Mining) to manage and conduct a Pre-feasibility Study on the viability of

mining the gold deposit from open pit mineral resources and processing ore at an annual

nominal production rate of 2.5 Mtpa to produce gold doré. The work has been undertaken in

collaboration with Amarillo and various other independent consultants, including Hoogvliet

Consulting Services (HCS) & Australian Exploration Field Services (AEFS), BVP Engenharia,

Onix Engenharia, Neotropica Tecnologia Ambienttal and Hidrovia Hidrogeologia e Meio

Ambiente.

2.1 Terms of Reference

This Independent Technical Report was prepared to provide technical information to support

the public disclosure of the Pre-Feasibility Study on the Mara Rosa Project of Amarillo. This

report was prepared in compliance with Canadian National Instrument 43-101 and provides a

summary of the full Pre-Feasibility Study report and Appendices that were prepared and

submitted to Amarillo in parallel with this report.

2.2 Qualified Persons

Qualified Persons responsible for the content of this technical report are:

Hugo Hoogvliet, MSc (Geology), MAusIMM, MAIG, Hoogvliet Contract Services; Geology QP;

responsible for Items 5 – 12. Since completing his contribution to this report Mr Hoogvliet has

passed away; as a consequence Mr Whitehouse, who is well familiar with the work

undertaken by Hoogvliet, has accepted sign off responsibility for these sections of the report.

Gregory Keith Whitehouse, BSc (Geology and Geography), MAusIMM (CP), Australian

Exploration Field Services Pty Ltd; Resources QP; responsible for Items 5 – 12 and 14.

Chris Witt, BSc Industrial Chemistry (Metallurgy), Diploma in Metallurgy, MAusIMM (CP),

Coffey Mining Pty Ltd; Mineral Processing and Metallurgical Testwork QP; responsible for

Item 13.

João Augusto Hilário, BSc (MinEng), MAIG, Coffey Consultoria e Serviços, Manager of

Technical Services and Mineral Reserve Estimates, Mining Methods, Project Infrastructure

QP; responsible for Items 15, 16 (excluding Item 16.1 and 16.2.5) & 18.

Clive Thomas Saunders, BSc, CGeol FGS,TMIE Aust, M.Zwe.I.E. Coffey Mining Principal

Tailings Engineer; prepared Item 16.2.5.

Frank Richard Baker, BMet, MMet, MIMMM, MAusIMM, Amarillo Gold Corporation; Amarillo

Project Manager and Recovery Methods QP; responsible for Items 17 and 19.



Norman Philip Lock, CGeol FGS, PrSciNat, Coffey Mining, Geology Manager; Study Manager

and QP responsible for Items 1 – 4, 16.1, 20 – 27.

Important study team members, not identified by QP status and whose work has been signed

off by one or other of the QPs above, include:

Adam Beer, Coffey Mining, Principal Geotechnical Engineer; prepared Item 16.1.

Francisco Laureano Fonseca, Coffey Consultoria e Serviços Ltda, Senior Mining Engineer;

prepared parts of Item 16.

Tetsue Oishe Coffey Consultoria e Serviços Ltda, Senior Mining Engineer; prepared parts of

Item 16.

Rod Smith, Salisbury Enterprises, Metallurgical Consultant; prepared Item 13.

Paulo Pessoa, Hidrovia Hidrogeologia e Meio Ambiente; prepared part of Item 20.2.2.

Janet Lowe, Coffey Geotechnics, Manager Environmental Services; prepared Item 20.

Amy Jacobsen, Senior Associate Consultant; prepared items 21 and 22.

BVP Engenharia; prepared Geotechnical Assessment report.

Neotropica Tecnologia Ambienttal; prepared Environmental Assessment.

Onix Engenharia e Consultoria Ltda; prepared Estudo de Pré-Viabilidade (Plant Design)

Elisabete Cançado Ferreira, Electrical Engineer

Hilton Souza Couto, Process Engineer

Ivan Antônio Dias, Mechanical Engineer

Renato Pinheiro, Civil Engineer

Curtis Clarke, VP Coffey Mining Canada; provided principal peer review

Jorge Schönherr, Associate Consultant Engineer; provided peer review of Mining Study (in

Portuguese).

Many other people in Belo Horizonte, Perth and Toronto have contributed in the preparation

of this report and these contributions are acknowledged here even though the individual may

not be named.



2.3 Site Visits and Scope of Personal Inspection

Hugo Hoogvliet has visited the Mara Rosa offices of Amarillo, the Posse mine site, the former

mine offices and core storage facilities on several occasions, most recently from 24 January

2011 to 31 January 2011. At the time of the visit holes MRP0033, MRP0034 and MRP0035

were being drilled. At the time of the visits a well-run team of geologists, field technicians and

a supervisor was active.

Paulo Pessoa of Hidrovia and Mariana Gomide of CERN carried out a site visit in the

company of Frank Baker of Amarillo from 28 to 31 March 2011 to inspect site conditions and

discuss hydrogeological study work.

Rod Smith and João Augusto Hilário of Coffey Mining carried out a site visit in the company of

Frank Baker of Amarillo and Hilton Couto of Onix from 4 to 7 April 2011 to inspect site

conditions, assess options for location of principal mine and engineering structures, discuss

and define battery limits between various study tasks, discuss and agree the process flow

sheet and to necessary additional metallurgical testwork.

Adam Beer of Coffey Mining carried out a site visit in the company of Frank Baker of Amarillo

from 31 July to 7 August 2011 to inspect the geotechnical drill program and progress, and to

discuss the core orientation and logging procedures. This visit was arranged opportunistically

for the ongoing feasibility study level geotechnical drill program but was of benefit to the

current PFS level study for a range of geotechnical issues from the mine pit, to general

infrastructure and tailings disposal site selection.

2.4 Effective Dates

The effective date of this report is taken to be the date of the finalization of inputs for the

financial model for the Project on 28 October, 2011. The dates for critical information used in

this report are:

The updated Mineral Resource estimate and mineral resource block model were

completed on 31 July, 2011

The Mineral Reserve estimate for the project was completed on 28 October, 2011

The final PFS mine plan was issued 28 October, 2011

PFS Mineral process engineering and capital cost estimation were completed 21 June,

2011

The PFS financial model was finalized 5 November, 2011

There were no material changes to the scientific and technical information on the Project

between the effective date and the signature date of the Report.



2.5 Information Sources and References

This Report is based on information provided in the following key documents and files:

Mineral Resource Block Model File - BM_MIK_AB_20110620_REP.XLSX (HCS & AEFS,

2011).

Report on Independent Site Visit and Resource Estimate. Posse Deposit, Mara Rosa, Goias

State, Brazil. (HCS & AEFS, 2011)

Geotechnical Report for Pre-feasibility Study. Mara Rosa Project (Beer, 2011)

Mine Schedule File – amarillo_sequenciamento.xlsx

PFS Report : Tailings Storage Facility and Water Storage Facility. Mara Rosa : Posse

Deposit. (Saunders, 2011).

Mara Rosa Metallurgical Testwork Report. Mara Rosa Pre-Feasibility Study. (Smith and Witt,

2011)

Estudo de Pré-Viabilidade. Mara Rosa Project. (Onix Engenharia, 2011)

Estudos Hidrogeológicos e Hidrológicos Básicos ‘PFS’ Mina De Posse, Mara / Go. (Hidrovia,

2011).

Environmental Assessment of the Mara Rosa Project Goiás State, Brazil. (Neotropica, 2011).

Capital Cost Estimate File - Capex_Projeto Posse_FL_Rev3.xlsx

Operating Cost Estimate File - Opex_ Mina_Posse.xlsx; Op Cost 03.10.Real 1.9 Energ. Exec.

Manning Excel Worksheet.xlsx

Financial Model File – Amarillo Gold Financial Model 07Nov11 DRAFT7.v2.xlsm

Coffey Mining has also sourced information from appropriate reference documents as cited in

the text and as summarized in Section 27 of this Report. Additional information was requested

from, and provided by, Amarillo. Coffey Mining has also relied upon other experts as outlined

in Section 3.

2.6 Units of Measure

Unless otherwise stated, the units of measure in this report are all metric in the International

System of Units (SI). All currency units are expressed in United States dollars (US$) or Brazil

Reais (R$), except where otherwise noted. Although some costs were derived from local

Brazil sources, all such numbers have been converted to US$ for presentation and financial

analysis. For Brazil costs an exchange rate of R$1.9 = US$1 was applied.



2.7 Previous Technical Reports

Amarillo has previously filed the following NI43-101 technical reports for the Project as

follows:

Caracle Creek International Consulting, 2008. Independent Technical Report and Preliminary

Economic Assessment, Mara Rosa Gold Property, Goias State, Brazil. Report prepared for

Amarillo Gold Corporation dated 29 February 2008.

Hoogvliet Contract Services and Australian Exploration Field Services PL. 2010. Independent

Mineral Resource Estimate and Preliminary Economic Assessment, Posse Deposit, Mara

Rosa, Goias State, Brazil. Report prepared for Amarillo Gold Corporation dated 30 June

2010.

Hoogvliet Contract Services and Australian Exploration Field Services PL. 2011. Report on

Independent Site Visit and Resource Estimate. Posse Deposit, Mara Rosa, Goias State,

Brazil. Report prepared for Amarillo Gold Corporation dated 30 July 2011.



3 RELIANCE ON OTHER EXPERTS

Neither Coffey Mining nor the authors of this report are qualified to provide comment on legal

issues associated with the Mara Rosa Project included in Section 4 of this report.

Assessment of these aspects has relied on information provided by Sabina and their legal

advisors, William Freire, Advocados Associados, and has not been independently verified by

Coffey Mining.

The source of this information is contained in:

William Freire. 2011. Synthetic analysis of the Exploration Permits referring the Mara

Rosa Project. Letter opinion on the legal status of mineral tenure by William

Freire, Advocados Associados, Belo Horizonte, MG, Brazil.

Coffey Mining has relied on the report of BVP Engenharia for geotechnical aspects of this

study. This report was reviewed by Coffey Mining and the abridged review presented in

Section 16.1 here has been supplemented by the additional discussion and reporting of Adam

Beer of Coffey mining and signed off by Coffey Mining.


BVP Engenharia, 2011. Visita Tecnica e Avaliacao dos Dados Geotecnicos do

Deposito Posse – Projeto Mara Rosa. Report prepared for Amarillo Gold

Corporation dated February 2011.

The authors of this report are not qualified to provide extensive comment on environmental

issues associated with the property referred to in Section 20 of this report. Assessment of this

aspect has relied heavily on a report by Brazil environmental company Neotropica Tecnologia

Ambienttal Ltda provided by Amarillo. This work has been reviewed by Coffey Mining for

compliance with Canadian reporting requirements.


Neotropica Tecnologia Ambienttal Ltda., 2010. Environmental Assessment of the Mara

Rosa Project Goiás State, Brazil, for Amarillo Mineraçao Do Brasil Ltda.

Neotropica Tecnologia Ambienttal Ltda., 2011. Environmental Assessment of the Mara

Rosa Project Goiás State, Brazil, for Amarillo Mineraçao Do Brasil Ltda.



4 PROPERTY DESCRIPTION AND LOCATION

4.1 General Description

The Mara Rosa Property (also generally known and referred to as the Posse Deposit) is

located in Goiás state, central Brazil, approximately 6 km north of the town of Mara Rosa. The

mining leases have an area of 2,552.62 ha and are centred at Latitude 13° 58.395′ S,

Longitude 49° 10.690′ W (approximate WGS84 coordinates 696880 mE, 8454530 mN, Zone

22 South) (Figure 4.1_1).

Western Mining Corp. (“WMC”) operated a small open pit mine at the project site during the

1990s. Two pits, Posse South and Posse North, were developed over a five year period. The

ore, along with feed from the nearby Zacarias mine, was processed on-site. The processing,

beginning with heap leach and later CIL, was conducted on approximately 10 ha of freehold

property adjacent to the mining leases. Local infrastructure included adequate power and

water to run a 600 tonne per day CIL plant and heap leach operation.

As of November, 2006, the mine and mill site had been reclaimed and no site infrastructure

remained. According to Amarillo, the required remediation for mine closure had been met and

accepted by the relevant government agencies. No significant environmental liabilities are

known to exist at the former mine site.

WMC maintained a core logging and storage facility, sample preparation laboratory, assay

laboratory, and office complex immediately north of the town of Mara Rosa. The facilities,

which occupy 8 ha of freehold land, have been utilized by Amarillo during their exploration

programs. As of January and April 2011 when various of the authors of this report visited the

project, the structures remain in excellent condition. The offices were utilized by Mr Hoogvliet

during his site visit and Amarillo staff during the drilling program which finished in March 2011,

Amarillo also owns two houses on contiguous pieces of land on São Paulo Street within the

town of Mara Rosa.



Figure 4.1_1 Location of Amarillo’s Mara Rosa Properties



4.2 Establishing Mineral Rights in Brazil

The process of acquiring title to mineral property in Brazil is a phased procedure involving

progressive categories of title as exploration and development work on a property advances.

Tenure is secure provided the title holder meets clearly defined obligations over time, but the

process of acquiring title can be lengthy.

Initially, an application for an exploration permit must be filed which meets specific regulatory

requirements and must include a description of the proposed exploration plan. A 60-day

period after filing is provided for the applicant to supply any additional information that may be

required. Then, if the applicant has met all requirements and if the area of interest is not

already covered by a pre-existing application or an existing exploration license, Exploration

licenses are issued by the Departamento Nacional de Produção Mineral (DNPM; the National

Department of Mineral Production), the Brazilian government mining agency.

Exploration licenses are granted for a period of three years, and are renewable upon request

for a further period of three years and are subject to a nominal charge per hectare

($R2.05/hectare). Exploration must begin within 60 days following issuing of the license, and

must be carried out according to the exploration plan. At the discretion of the DNPM,

exploration licenses may be terminated if exploration activities are suspended for more than

three years. An applicant may reapply for a terminated license at the end of the three-year

license period. The reapplication must include a new exploration plan and the re-issued

license is subject to an increased annual fee.

Mining licenses are only granted to corporations. Normally, applicants have a period of one

year, following DNPM approval of the final exploration report, on an exploration licence, to

present a mining plan and a feasibility study, and to apply for a mining license. No fees are

levied on the holder of a mining license and a mining license does not convey title to a mineral

deposit which remains vested in the government. Rather, a mining license gives to the holder

the right to extract, process, and sell minerals from the deposit, in accordance with a plan

approved by the DNPM, until the deposit is exhausted.

Presently, Amarillo holds a property position totalling 80,833.9 ha of exploration leases and

2,552.62 ha of mining leases centred over the Posse gold mine (Figure 4.2_1). The mining

leases have been surveyed as required by Brazilian mining law; however, the exploration

leases have not as they are recorded on a graticular system. The list of mining claims over

and around the Posse Deposit, as registered with DNPM on 26/1/2011, is presented in Table

4.2_1.

While the granting of an exploration concession gives Amarillo the required permits for

conducting exploration activities on the property, very limited work is currently being carried

out on the exploration concessions, as all of the company’s resources are dedicated towards

a production decision on Posse. The company is negotiating with the DNPM in good faith to

retain the exploration leases that are up for renewal on the basis that its resources are



committed to the task of establishing a mine at Posse. Once the decision on Posse has been

made, the focus will return to the regional work.

Figure 4.2_1

Mining and Exploration Concessions over and around the Posse Deposit



Table 4.2_1

Mara Rosa Project

Mara Rosa Mining Concessions

Tenement ID Area (ha) Concession Type

860952/1980 1000.00 Mining Concession



As the Posse mine is no longer in production, the mining concessions have been registered

as under “Mining Suspension”, a status allowing them to be held without production occurring.

Amarillo recently lodged a Reserve Evaluation Report with the last works developed by the

Company and requesting three more years of extension of the “Mining Suspension”. In a

meeting with the DNPM´s superintendent and inspectors they confirmed, verbally, that the

three years will be granted, but some exigencies should to be accomplished up to end

December. The activities requested by DNPM in the exigencies already are being carry out

by Amarillo. To support the claim for extension Amarillo have provided DNPM with details of

the Mara Rosa mineral resources, its relationship to the Mining Concession boundaries and

other relevant information.

A positive legal opinion on Amarillo’s mineral tenure has been provided by William Freire,

Advocados Associados (2011), with some comments on a timeframe for conditional

compliance.

“Based on the information in the mining registry, we conclude that proceedings 860.952/1980,

861.241/1980 are 862.000/1984 are, at this time, regular.

This regular status should be understood to mean the following:

a) There is no administrative or judicial proceeding that seeks to obtain a declaration of

caducity or nullity or to transfer ownership of the Mining Titles.

b) The Mining Titles do not guarantee compliance with other obligations.

c) The administrative proceedings follow the procedure set forth in the mining code and

related legislation with regards to the timeliness of the reports and other exclusively

technical documents necessary to maintain the Mining Titles.

However, there is:

a) A compliance restriction in relation to the Annual Mining Reports for base years

2000, 2002, 2003, 2004, 2009 and 2010.

b) A need to comply with two immediate obligations.”



Freire concludes the immediate obligations are:

Presentation of Annual Mining Report for base year 2011 by 15 March 2012; and

Compliance with the requirements published on 27 October 2011.

These should be concluded within the “established timeframe”.

4.3 Royalties and other agreements on the property

Royal Gold has acquired, through the purchase of Metallica, a 1% NSR Royalty.

Franco Nevada has acquired a 1% NSR Royalty from Euro-Nevada Gold Corporation.

4.4 Environmental Liabilities

There are no current environmental liabilities known for the property.

4.5 Permits required for Development

(g) to the extent known, the permits that must be acquired to conduct the work proposed

for the property, and if the permits have been obtained; and

4.6 Other Factors or Risks affecting Access, Title or ability to Work

(h) to the extent known, any other significant factors and risks that may affect access, title,

or the right or ability to perform work on the property.

In the opinion of Coffey Mining, HCS and AEFS, Amarillo have the tenements under suitable

tenure to enable the future development of the Mara Rosa project.



5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

Details of the of surface rights, availability of power, water, labour and both waste disposal

and process plant locations are discussed in more detail in separate relevant sections of this

report. Suffice it to note that the project is not seriously constrained by space or other factors

necessary for mining activities.

5.1 Accessibility

The Municipality of Mara Rosa is located 356 km north of Goiânia in the Porangatu

Microregion, 11 kilometres west of the Belém-Brasília highway, between the basins of the

Araguaia River and the Tocantins River. According to a 2005 estimate, Mara Rosa has a

population of approximately 25,000 people of whom 11,500 live in the town.

5.2 Climate

Average annual rainfall is 1,679 mm (as measured at Estrela de Norte 30 km to the north),

resulting in a relatively wet climate. The year is defined by two principal seasons, a dry

season from April to September and a wet season from October to March. The mean

temperature is 24 °C during the dry season and 28 °C during the wet season. Annual

temperatures typically range from approximately 4 °C to 45 °C. The climate does not impact

operations which continue throughout the year.

5.3 Local Resources

Local facilities include several public and private elementary and high schools, two hospitals,

a public health centre, one bank, two small motels and numerous shops. Agriculture (corn,

rice, manioc, sugarcane, soybeans, and bananas) and cattle ranching are the primary

commercial activities in the region; Mara Rosa is a regional support community for these

activities.

5.4 Infrastructure

The municipality has an excellent network of local farm roads, the majority of which are

unpaved but in generally good condition. The municipality is also serviced by an 800 metre-

long, unpaved airstrip. Access to Mara Rosa is via Federal Highway BR-153, the main north-

south highway in central Brazil leading to the city of Belém at the mouth of the Tocantins river.

Mara Rosa is 356 km, or 4 hours driving time, north from the state capital of Goiânia, and 320

km, or 4 hours driving time from the national capital, Brasilia. Highway communications

(Figure 5.4_1) with Goiânia are made by GO-080 / Nerópolis / São Francisco de Goiás / BR-

153 / Jaraguá / GO-080 / Goianésia / Barro Alto / GO-342 / BR-153 / Uruaçu / Campinorte /

GO-239.

Electric power is supplied by CELG, the State of Goiás Energy Authority. The local electricity

grid is rated at 69 kW. The water supply is metered and is provided by SANEAGO, the state



water company. Water for the Posse mine site as well as ranches in the surrounding region is

derived from a combination of local streams and artesian wells. Telephone service, both local

and international, is provided by TELEGOIAS. Cellular telephone service is available in the

area.

Figure 5.4_1

Mara Rosa, surrounding towns



5.5 Physiography

The region is characterized by tropical savannah of low to moderate topographic relief ranging

from approximately 400 m to 500 m above sea level (ASL). The town itself has a mean

elevation of 520 m ASL. Much of the area has been cleared for farming and as a result is

open savannah grassland. Trees occur along the abundant water courses.



6 HISTORY

This section summarises the work carried out prior to the 2010 - 2011 drilling campaign. The

work carried out from 2010 is discussed in Section 10.

Evidence of small scale surficial-alluvial mining along the Rio do Ouro in the historic Amaro

Leite area indicates mining activity in the Mara Rosa District dates to the mid-1700s. More

recent activity dating from the early 1970s to early 1980s began with the successful

discoveries by INCO (now Vale S.A. or “Vale”) of the Chapada gold-copper and Crixás gold

deposits. These deposits are located approximately 30 km and 100 km to the south-west of

the town of Mara Rosa, respectively.

During the early 1980s, BHP-Utah Mines (now BHP Billiton Limited), through its subsidiary

Mineração Colorado Ltda., initiated a grass roots reconnaissance program that covered the

Chapada district and the Mara Rosa area, and eventually led to the discovery of the Posse

gold and Zacarias gold-silver-barite deposits. From 1981 to 1987, BHP completed 12,300 m

of diamond and reverse circulation (RC) drilling at Posse and Zacarias. At Posse, a 107 m

exploration shaft was sunk and 400 m of lateral drifting was completed to test mineralization.

As a result of Brazilian restrictions on foreign ownership at the time, in 1988BHP chose to

joint venture the Mara Rosa properties with Western Mining Corp. In 1990, WMC set up a

subsidiary, Mineração Jenipapo S.A. (“MJSA”), to acquire a 100% interest in Posse, and to

explore, develop, and operate the asset. The Posse mine was opened in 1992 and operated

until July, 1995 during which time two pits, Posse North and Posse South, were developed.

The on-site mill processed approximately 750,000 tonnes of ore grading a combined 3.5 g/t

Au. Zacarias, which was significantly higher grade, operated at roughly the same time as

Posse and was processed through the same mill.

In order to provide cash flow for its activities in Brazil, WMC focussed much of its attention on

development of the Posse and Zacarias mines between 1990 and 1995. This work is

understood to have been completed as a result of a corporate decision to make each

business unit self-funding and to encourage efforts to develop known deposits. In addition,

efforts to replace mined reserves were directed toward both the Eastern and Central Belt

exploration targets generated previously by BHP as well as new targets identified to the east

and north of Mara Rosa.

By June 1995, a combination of factors, including low gold prices, the exhaustion of reserves

at the higher grade Zacarias deposit, and the failure to discover any additional, near-surface

reserves, caused WMC to discontinue mining and exploration activities at Mara Rosa. As the

primary exploration objective had been the discovery of near-surface mineralization that could

be fast-tracked into production, most of the exploration targets identified by BHP and WMC

had only been evaluated to depths within, approximately, 50 m from the surface.

Upon suspension of its mining and exploration activities, WMC was approached by several

companies interested in exploring the property under lease-option agreements. The Zacarias



deposit and the rights to its tailings were eventually sold to Minere Mineração Ltda. (“Minere”),

a small Brazilian company interested in exploiting the deposit’s very high barite content. The

project has since been on-sold to a company called Baribras Mineração Ltda.

In 1996, Barrick do Brasil (“Barrick”) completed a full due diligence study of the remaining

Posse project concessions (the Eastern Belt claims). The due diligence involved a team of at

least 14 people and a significant program of test sampling, re-logging of core, soil sampling,

reinterpretation of geophysics, and an estimate of the mineral resource for the Posse Deposit.

Although this program subsequently led to a preliminary offer by Barrick to purchase the

property in full, negotiations stalled prior to execution of the agreement. Barrick provided

WMC with a copy of its due diligence report and related correspondence after the failure to

execute a deal.

Following Barrick’s withdrawal, Metallica entered into negotiations with WMC for purchase of

the Eastern Belt properties, and in November, 1997, successfully completed an agreement

which called for a total purchase price of US$1.5 million. As part of the previous buy-out

agreement between BHP and WMC, BHP held a 1% NSR royalty interest on the property.

This now sits with Royal Gold and, as a part of the Metallica purchase, Euro-Nevada Gold

Corporation (later absorbed into Newmont) held an additional 1% NSR royalty. This now sits

with Franco Nevada Corporation.

Following a compilation of data and a review of the project, Metallica completed a systematic

soil geochemistry and geological mapping program north-east of the Posse Deposit. Induced

Polarization (IP) and ground magnetic geophysical surveys were completed over some of the

more promising areas. Metallica suspended exploration operations in September, 1998, and

placed the project on care and maintenance. In 2001, Metallica revisited the project and

completed a review of the regional potential. At this time, 5 holes, totalling 940 m, were drilled

into three separate targets on the northern extensions to the Posse mine trend. Following this

work, a corporate decision was made to focus on properties in Mexico and Chile and Metallica

decided to sell the project.

Amarillo Gold Corporation (Amarillo) visited the project in August, 2003 and in October, 2003

signed a letter of intent with Metallica to purchase MBL and 100% of the Mara Rosa project.

The project remains subject to the 1.0% NSR royalty to Franco Nevada Corporation and a

further 1.0% royalty to Royal Gold.

6.1 Exploration History

BHP held the Property from 1982 until 1987; WMC from 1988 until 1997; and Metallica from

1997 to 2003. The site has been controlled by Amarillo since then. The following is a

summary of work known to have been completed on the Property prior to the release of the

NI43-101 Technical Report, dated July 2011 for the property:

244 exploration drill holes totalling approximately 25,070.7 metres (HCS & AEFS 2010),

these are further discussed in 6.4 below;



4,011 blast holes with corresponding assay data for 6,598 samples (Hoogvliet, 2008);

these holes were drilled as part of the WMC mining operation on the Posse North and

South Pits.

307.5 metres of underground sampling recorded as 10 horizontal drillholes;

278 metres of surface channel sampling;

Geochemical sample data, Metallica verification sampling program, 2,947 samples;

Multi-element lithogeochemical data (575 Posse core samples and 31 regional samples);

3D geological model and drill section plots used for the 1997 Posse resource estimate;

3D geological model and drill section plots used for the 2008 Posse resource estimate by

CCIC;

3D geological model and drill section plots used for the 2010 Posse resource estimate by

HCS & AEFS;

Posse pit geology map at 1:50 scale (Metallica mapping, AutoCAD file);

North Posse trend district geologic map at 1:5,000 scale (Metallica mapping, AutoCAD

file);

Eastern Belt regional geologic map compiled at 1:10,000 scale (Metallica and BHP

mapping, AutoCAD file);

Grid E ground magnetics data, unedited BHP data (Excel spreadsheet).

In addition, a total of thirty detailed geologic and alteration sections were constructed by

Metallica for the Posse Deposit, these remain to be digitized into the project database but will

be superseded by work currently being undertaken by Amarillo to update formal geological

cross sections

6.2 Metallica Exploration

Metallica held the Mara Rosa project from 1997 until 2003. Between July 1997 and January

1999, Metallica focussed on the development of a detailed geologic model for the Posse

Deposit, and on the verification and improvement of exploration targets along its general

strike extensions. No drilling was completed by Metallica during this time. Exploration work

completed by Metallica included the following:

Digital compilation of geological, geochemical, and geophysical information from the

project data files;

Geological mapping and structural analysis of Posse pit and re-logging of approximately

8,100 m of Posse drill core;



Systematic lithogeochemical sampling through the Posse Deposit and Eastern Belt

stratigraphic sequence;

Geological mapping (1:5,000 scale) and soil sampling along approximately 23 km of

Eastern Belt/Posse strike extensions;

Completion of detailed IP-resistivity and ground magnetics surveys over Posse and its

strike extensions;

Regional reconnaissance of the Far Eastern, Central, and Western Belt areas.

In January, 1999, due to the bear market in gold and Metallica’s corporate decision to focus

on other projects in Chile and Mexico, the Posse Deposit was placed on care and

maintenance. A soil sampling campaign was completed in 2001. In late 2002, with the

improving gold price, a 5 hole drill program was completed on three separate targets along

the northern strike extent of the Posse Deposit. Following this work, and the completion of a

resource estimate of the Posse Deposit, Metallica decided to sell the Mara Rosa project.

6.3 Amarillo Exploration

Amarillo Gold Corporation first visited the Mara Rosa project in August, 2003 and

subsequently signed a letter of intent with Metallica to purchase MBL and 100% of the Mara

Rosa project in October, 2003. Micon International Ltd. was retained by Amarillo to review the

Property in late 2003. and B. Terrence Hennessey visited the Property November 14 to 17,

2003 and is the sole author of the December 2003 report: A Review of the Mara Rosa Gold

Project, Goias State, Brazil (summary published on www.sedar.com January 13th, 2004 as

part of the Amarillo 2003 annual report).

In 2005 further work was undertaken to verify the location of previous drilling and generate a

volume of waste dumps. This was followed, in 2006, by a trenching program of 28 trenches

totalling 2,942 metres to verify previous soil sampling. Results from the trenching program

were used to sight some of a series of 28 diamond holes (SPETI-01 to SPETI-28). The

SPETI drill program ran from 1 December 2005 until 26 September 2006. Results from the

SPETI program together with data extracted from the historical drilling database were used by

CCIC to generate a Resource Estimate. An updated Resource Estimate for the property

together with a Preliminary Economic Assessment was contained in an NI43-101 Technical

Report dated 29 February 2008 (published on www.sedar.com 26 March 2008). Further work

was undertaken in 2008. This consisted of 14 diamond drillholes totalling 164.7 m. This

program was interrupted by the GFC. At the same time a recompilation of the geological

database was undertaken by HCS. This (recompiled) database was used as the data source

for the updated resource estimate and Preliminary Economic Assessment contained in a

NI43-101 Technical Report dated 30 June 2010 (published on July 12, 2010 at

www.sedar.com.



6.3.1 Validation of Drill Hole Locations

In 2005, Kevin R. Kivi, P. Geo, acting as consultant to Amarillo recommended that an

accurate GPS survey of historical survey locations be undertaken to allow all data to be

recorded in UTM coordinates based on the WGS84 Datum as previous drilling had been tied

to survey monuments located in terms of local grids. Subsequently GLOBAL Informática e

Consultoria was contracted by Amarillo to complete a differential GPS survey of available

landmarks on the Property and UTM coordinates based on the WGS84 datum were

established for most historical drill collars.

6.3.2 Waste Dump Volume

As a part of the work carried out by GLOBAL Informática e Consultoria, 3,275 survey points

were taken on the waste pile in order to provide data for a volumetric calculation. This data

was utilised by CCIC to create a digital elevation model (DEM) of the waste pile and the

volume of this DEM was calculated to be 321,800 m3.

6.3.3 Surface Trenching

During 2006, Amarillo excavated 28 trenches generally between 2 and 3 m in depth totalling

2,942 m. The purpose of the trenching was to evaluate the better geochemical anomalies

identified by Metallica’s extensive soil sampling program. Following systematic 1:100 scale

mapping and logging of each trench samples were collected along the length of the trenches

at 2.5 m intervals.

The degree of metamorphism and weathering was reported to present a significant challenge

to identifying lithology however some features such as quartz veining, pegmatities and

dolerite dykes were readily identifiable in the trenches.

The gold values reported by Metallica in the soils were generally confirmed, but the values

returned from analysis of the trenched, weathered rock were often no higher than the values

in soils.

6.4 Historical Drilling

A number of drilling campaigns have been completed on the Mara Rosa tenements. These

have been carried out by:

BHP from 1982 until 1987,

WMC from 1988 until 1995,

Metallica in 2002

Amarillo in late 2005 and 2006

Amarillo in 2008



In all, 233 drill holes were completed by BHP and WMC for a total of 19,876.7 m between

1982 and 1995 (Hennessey, 2003). The drilling concentrated on the Posse Deposit.

Subsequently Metallica drilled five drill holes and Amarillo a further 42 drill holes.

BHP conducted two main campaigns of drilling. An initial round of percussion drilling (hole

numbers beginning with W prefix), followed by a second round of diamond drill core holes.

The percussion drill holes largely targeted potential for a shallow oxide resource and were

drilled on a 50 m by 50 m grid. The diamond drill holes targeted potential for a deeper

sulphide resource and were also drilled on a 50 m by 50 m grid mainly over the Posse

Deposit to a depth of 350 m. At the periphery of mineralization, drill hole spacing was widened

in multiples of 50 m.

Following this work, WMC completed 107 reverse circulation (RC) and diamond drill holes

which mostly targeted the Posse Deposit. The RC drill holes were shallow and drilled on a 25

m by 25 m grid pattern. The diamond drill holes were deeper and followed on from the BHP

drilling of the sulphide resource.

The completed drilling was utilized to estimate the ore reserves mined from the Posse North

and South pits as well as the remaining historical mineral reserves and / or resources

previously reported by WMC, Barrick, and Metallica. Although both WMC and BHP completed

some shallow percussion style drilling on a few exploration targets, minimal drilling was

completed distal of the main Posse Deposit area.

In November 2002 Metallica completed five diamond drill holes totalling 940 m on three

separate targets along the Posse trend. The targets tested were an inferred structural plunge

projection of the Posse North lode and two geochemical-geophysical anomalies located along

strike from the Pose deposit.

Between December 1st, 2005 and September 27th, 2006, Amarillo completed a diamond

drilling campaign that totalled 3,461.2 metres in 28 holes (SPETI-01 to SPETI-28). This

drilling was completed utilizing a MACH FS 320 (serial number 922) drill rig constructed by

Maquesonda of Rio de Janeiro, Brazil in April, 2005. The objective of Amarillo’s drill campaign

was to confirm and expand known limits of gold mineralization within the Posse Deposit, as

well as to test previously unevaluated exploration targets identified by predecessors. Drill

targeting was based on previous data and results from 26 trenches completed by Amarillo in

2005.

The drill holes were sunk through the saproltic overburden into fresh bedrock using HWL

casing; followed by NQ diameter core (47.6 mm) to total depth. Fresh bedrock was generally

intersected at a true depth of 5 to 15 metres. Each run of the MACH FS 320 is approximately

3 metres and the average core recovery in each hole was reported to range from 70% to

80%.



The drilling outlined above was used for the Resource Estimate and Preliminary Economic

Assessment contained in the NI43-101 Technical Report dated 29 February 2008 prepared by

CCIC.

In 2008 Amarillo conducted a further diamond drilling campaign commencing May 21 2008

and finishing October 16, 2008. 15 holes totalling 3,428.7m were drilled. Hole_ID's were

W002A and MRP0001 to MRP0014 and the drilling was completed by Brazilian company

Servitec Sondagem Geologica using a Maquesonda 1200 drill rig using standard triple tube

wireline drilling.

The total drilling completed by BHP, WMC, Metallica and Amarillo is summarized in Table

6.4_1 (Drill hole information prior to 2003 is from Hennessey, 2003).

Table 6.4_1

Mara Rosa Project

Summary of Historic Drilling

Company Drilling Meters

BHP 54 percussion holes 3,635.3

72 diamond drill holes 10,902.3

WMC 57 RC holes 2,060.3


Metallica 5 diamond holes 940.0

Amarillo 43 diamond holes 6,802.4

Total 266 holes 27,619.1

The drillhole data included in the Posse database at the end of 2009, which was used for the

2010 resource estimation, comprised the holes listed in table 6.4_2.

Table 6.4_2

Mara Rosa Project

Summary of Drillholes in the Posse database end 2009

Company Drilling Meters

BHP 34 percussion holes 2590.4

45 diamond drill holes 2930,2

WMC 53 RC holes 2,040.0


10 underground drillholes 307.5

Amarillo 43 diamond holes 6,802.4

Total 244 holes 25,075.7

There is a discrepancy between the number of exploration holes reported as drilled and the

number of holes in the geological database for Posse. The discrepancies are in holes related



to pre Amarillo drilling as the exploration holes summarized above include holes drilled

outside the area covered by the Posse dataset. In addition it would appear that the Metallica

holes have been renumbered and now appear in the Posse database as being WMC holes. It

is recommended that further work be done to reconcile these differences, however it is the

opinion of HCS and AEFS that the issue does not pose a material threat to the integrity of the

Posse database or the Resource reported in this technical report.

6.5 Historical Resource Estimates

Five mineral resource estimates have been prepared for the Posse Deposit subsequent to

mine closure. These are set out in Table 6.5_1 below.

Table 6.5_1

Mara Rosa Project

Historic Grade Tonnage Estimates

Report Comment

WMC,1995 In house estimate, immediately after mine closure *

Barrick, 1996 Estimate completed as part of its 1996 due diligence study *

Metallica, 1997 Estimate, completed prior to sale of project *

Amarillo, 2008 Undertaken by CCIC, NI43 - 101 compliant.

Amarillo, 2010 Undertaken by HCS & AEFS, NI43 - 101 compliant.

* These resource estimates are not considered compliant to the terms set out in NI43-101 and

are referenced for historic purposes, they should not be relied upon. A qualified person has

not done sufficient work to classify the historical estimates as current mineral resources; and

Amarillo is not treating these historical estimates as current mineral resources or mineral

reserves.

6.5.1 WMC Grade Tonnage Estimate

According to the Barrick due diligence documentation, WMC reported, at the time of mine

closure, a remaining open-pitable reserve of 1,163,433 tonnes grading 2.08 g/t Au. This

“reserve” was reported at a cutoff grade of 0.6 g/t Au and a stripping ratio of 2.3:1; no

reference to categorization was given. This reserve was limited to 100 m below the surface.

WMC documentation pertaining to this estimate has not been viewed by the authors. Coffey

Mining notes that this estimate cannot be relied upon for the reasons stated in Table 6.5_1.

6.5.2 Barrack Grade Tonnage Estimate

As part of its due diligence, when Barrack was considering purchase of the property, Barrick

arranged for the completion of a mineral resource estimate for Posse South after confirmation

of the geological model by the due diligence team. The estimate was completed by Pedro

Guzman of PGV Consultores in Chile. Barrick’s resource was reported at a cutoff grade of

1.0g/t Au and to a depth of 300 m below surface. The result of the mineral resource estimate,

estimated by Ordinary Kriging, was 15,468,300 tonnes grading 1.47 g/t Au. No confidence



categories were assigned to the estimate. Coffey Mining notes that this estimate cannot be

relied upon for the reasons stated in Table 6.5_1.

6.5.3 Metallica Grade Tonnage Estimate

In late 1997, Metallica estimated a mineral resource for Posse South after re-logging

approximately 8,100m of core and remapping exposures in the pit. Shallow reverse circulation

and blast holes from the pit were also used. The estimate was completed using Ordinary

Kriging, a 10.0 g/t Au top cut, and a cutoff grade of 1.0 g/t Au. Metallica’s resource estimate

was also prepared by Pedro Guzman from a new geological model interpreted by Metallica

staff. The Metallica resource estimate is separated into the measured, indicated, and inferred

confidence categories (Table 6.5.3_1). However, Mr Guzman’s report does not state which

resource reporting code was used. Coffey Mining notes that this estimate cannot be relied

upon for the reasons stated in Table 6.5_1.

Table 6.5.3_1

Mara Rosa Project

Metallica Resource estimate (1.0 g/t Au cutoff)

Category Tonnes Grade (g/t Au) Contained Ounces

Measured 6,385,000 1.81 372,300

Indicated 4,948,000 1.67 265,700

Inferred 1,417,000 1.94 88,500

Total 12,750,000 1.77 726,500

6.5.4 Amarillo; CCIC Resource Estimate

CCIC was retained by Amarillo to complete an Independent Mineral Resource Estimate and

Preliminary Economic Assessment of the gold resources located on the Mara Rosa Property,

and to produce a supporting Technical Report in accordance with the guidelines set out in NI-

43-101, companion policy NI43-101CP and Form 43-101F1. CCIC completed an initial

assessment of an Inferred Mineral Resource Estimate in March 2007 and an updated mineral

resource estimate together with a preliminary economic assessment was published in

February 2008.

This mineral resource estimate (Table 6.5.4_1) was assessed using Ordinary Kriging, with a

70 m x 50 m search ellipse and top cutting of gold grades at 30 g/t Au.



Table 6.5.4_1

Mara Rosa Project

Posse Deposit mineral resource estimate

Category Tonnes Grade (g Au/t) Contained Ounces

Indicated 13,515,000 1.48 643,000

Inferred 13,326,000 1.26 538,000

In addition to the resource estimate, CCIC completed a Preliminary Economic Assessment of

the Posse Deposit based on the results of a Lerch-Grossman pit optimization carried out

using Gemcom’s Whittle software based on a base gold price of US$800/oz Au (US$25.72/g

Au).

With the parameters utilized in the CCIC PEA, the break-even gold price for the project was

approximately US$575/oz. At higher gold prices, the project value increased significantly. At

US$600/oz, the project returned approximately US$12 Million. At US$900/oz (the then market

conditions), the project returns approximately US$153 Million. At US$1,500/oz, the project

returns approximately US$485 Million.

6.5.5 Amarillo; HCS & AEFS Resource Estimate 2010

In 2010 HCS and AEFS were retained by Amarillo to produce a Resource Estimate and

Preliminary Economic Assessment of the Posse Deposit. The work was the subject of an

NI43 - 101 technical report published on SEDAR July 12 2010. The resource model was

constructed using Median Indicator Kriging with a 70 m x 35 m x 7 m primary search ellipsoid

to populate a 25 m x 25 m x 10 m block model. Blocks were allowed to be sub-blocked to 5 m

X 5 m X 5 m to fit the constraining mineralisation wireframes. At a 0.5 g/t cutoff the resource

was as shown in table 6.5.5_1.

Table 6.5.5_1

Mara Rosa Project

HCS & AEFS Resource Estimate 2010

Cutoff Category Tonnes Grade Contained Oz

0.5 g/t Indicated 11,928,000 1.62 623,000

Inferred 10,164,000 1.38 451,000

In addition to the Resource Estimate, a Preliminary Economic Assessment was made based

on the results of a Lerch-Grossman pit optimization carried out using Micromine software

based on a gold price of US$1,000/oz Au (US$32.15/g Au).



The breakeven gold price was approximately US$475/oz Au with an optimal pit taken to be

the "US$850" pit shell reached after 12 years of mining. This pit shell has an NPV of

USD197m and would see a recovery of 800,000 oz of gold at a price of US$1,000/oz.

6.6 Historical Production

The Posse mill and leach pads processed ore from two deposits, Zacarias and Posse,

between 1992 and 1995. Until the end of June, 1995 production from the Posse North and

South pits totalled approximately 388,000 tonnes (Table 6.6_1). Mining continued for six days

into July, 1995, however, records for this final part month of production have not been located.

Records from the oxide heap leach have not been located.

Table 6.6_1

Mara Rosa Project

Summary of WMC Production at Mara Rosa, Posse Deposit Sulphides

Extraction Method Tonnes Grade (g/t Au) Contained Ounces

CIL 236,356 3.01 22,873

Heap Leach 152,103 1.62 7,922

Total 388,459 2.47 30,795



7 GEOLOGICAL SETTING AND MINERALISATION

7.1 Regional Geology

The Mara Rosa District is situated within the Goiás Magmatic Arc (GMA) which forms part of

the Tocantins physiographic province, an intercratonic mobile belt that separates the

Amazonas and São Francisco cratons, located to the north-west and south-east respectively

(Figure 7.1_1). The GMA is a 100 km wide, northeast-trending granite-greenstone terrane that

extends approximately 700 km. It is composed of accreted Middle to Neoproterozoic volcano-

sedimentary island arc sequences that have been intruded by granitoid and mafic plutons

(Kuyumjian, 1998). Subsequent metamorphism of these units has resulted in their

recrystallization to upper greenschist to epidote-amphibolite facies.

The geology in the Mara Rosa District is principally delineated by three northeast-striking,

moderately to steeply northwest-dipping belts of metamorphosed volcano-sedimentary and

associated intrusive rocks. These belts, referred to as the Western, Central, and Eastern Belts

(Figure 7.1_2), are separated by broad zones of tonalitic orthogneiss. The Eastern Belt is

bounded to the southeast by the Rio dos Bois fault, which also defines the south-eastern limit

of the GMA. The Rio dos Bois fault is a major regional fault that has thrust Neoproterozoic

rocks of the GMA over older, Early to Middle-Proterozoic rocks of the Central Goiás Massif

positioned to the south and east. Structures within the upper plate of the Rio dos Bois fault

(i.e. the Central-Eastern shear zone) are understood to be the principal control for

mineralization in the Mara Rosa region.

Several significant mineral deposits occur in the Mara Rosa District. These include the Posse

gold deposit, the Zacarias gold-silver-barite deposit, and the Chapada copper-gold deposit. A

relatively widespread distribution of potential gold, copper, and silver mineralization is

suggested by the numerous historic prospects (garimpos) that are dispersed in the regions

between these deposits. The Crixás gold district, located approximately 100 km to the south-

west of Mara Rosa, lies immediately outside the mapped limits of the GMA.



Figure 7.1_1

Summary geology of Brazil



Figure 7.1_2

Mara Rosa Local Geology



7.2 Local Geology

Amarillo’s land position within the Mara Rosa District primarily covers the Eastern Belt

greenstone assemblage (Figure 7.1_2). As a result of WMC’s sale of the Zacarias deposit to

Minere, Amarillo’s land position includes limited ground covering the Central and Western

Belts.

The Eastern Belt, which in general strikes to the north-east and dips moderately to steeply to

the north-west, has a maximum thickness of 6 km. Surface topography over the belt is

characterized by moderate relief and locally dissected drainages that follow lithologic or

structural weaknesses. Depth to fresh bedrock is generally shallow, ranging from 0 to 15 m.

The upper portion of the weathered profile consists of clay-rich latosol and saprolite derived

from the underlying bedrock. Low-lying areas along drainages channels are covered by a

relatively thin (<15 to 25 m) veneer of alluvial sediment. Although fresh outcrop is sparse,

individual stratigraphic units can be distinguished by surface floats and local soil/saprolite

composition. The principal stratigraphic units of the Eastern Belt are presented in Table 7.2_1

and Figure 7.2_1.

Table 7.2_1

Mara RosaProject

Principal stratigraphic units of the Eastern Belt

Rock Unit Lithology Description

Western Amphibolite Moderately foliated hornblende-plagioclase-quartz amphibolite

Grey Gneiss Moderately foliated feldspar-biotite-quartz gneiss

Posse Schist Well foliated quartz-biotite-sericite schist with feldspar (sheared product of Grey Gneiss)

Medial Metagabbro Strongly foliated hornblende-plagioclase amphibolite with biotite

Muscovite Schist Strongly foliated muscovite-quartz & ankerite-pyrite schist

Eastern Amphibolite Well foliated/laminated hornblende-plagioclase & biotite-quartz-ankerite amphibolite

Amaro Leite Sediments Well foliated fine-grained feldspathic quartz arenites plus minor interlayered amphibolites and ferruginous cherts



Figure 7.2_1

Mara Rosa District stratigraphic column

These rocks of the Eastern Belt are locally intruded by quartz-feldspar-muscovite & biotite

granitoid bodies and associated aplite and pegmatite dykes, small stocks and dykes of

hornblende & biotite & magnetite diorites, and, in its north-central portion, a large body of

hornblende-plagioclase gabbro. All units exhibit varying degrees of foliation that typically

range from weak to moderate, and generally intensify along sheared contacts. The tonalitic

orthogneiss which separates the Eastern and Central Belts is composed of coarse-grained

plagioclase, hornblende, and biotite with localized patches of biotite schist near its contact

with the Eastern Belt.



The Eastern Belt is structurally dominated by well-developed, penetrative foliation that strikes

30° to 50° and dips 40° to 70° north-west – an orientation subparallel to stratigraphy. Major

structural systems include 50° to 65° striking shears and thrusts and associated drag folds.

Shears are most commonly developed along zones of elastic disparity such as lithologic

contacts. Shear sense is typically reverse dextral oblique although a sinistral sense is locally

observed. A second set of structures consists of late cross cutting north-west to east-

northeast striking brittle faults and fractures that locally offset stratigraphy in apparent dextral

strike-slip sense. Although the style of deformation for the region is strongly suggestive of

isoclinal folding and possible stratigraphic inversion, direct field evidence indicating the

presence of such phenomena remains inconclusive (Metallica personal communication to P.

Mullens).

Uranium-lead isotopic age determination of zircons from some of the principal lithologic units

within the district indicates timing of initial rock formation for both the belt rocks and the

tonalite gneiss to be between approximately 870 Ma to 850 Ma (Viana, 1995). Subsequent

amphibolite facies metamorphism is estimated to have occurred between 700 Ma to 600 Ma

based on uranium-lead and rubidium-strontium dating of recrystallised titanate. The latter date

corresponds to peak metamorphism related to the Brasiliano orogenic event.

7.3 Property Geology

The mineralisation envelope at Mara Rosa is about 30 m thick and over 1 km long (Figure

7.3_1). It is structurally controlled dipping to the NW at about 45°, striking NE-SW. The

mylonitic appearance is most noticeable in the footwall where shearing is the most intense.

The intense shearing is associated with generally increased sulphide mineralisation (up to

about 4%), and a slight increase in metamorphic grade from greenschist to high greenschist

facies in the hanging wall to high greenschist/low amphibolite facies in the footwall (large

biotite flakes, some garnets). The best gold values are associated with intense shearing and

higher levels of sulphide mineralisation.

Aside from the slight increase in metamorphic grade, there appears to be a chemical

difference between the hanging wall and footwall, based on ICP analyses obtained from the

2005/2006 drilling program, this however cannot be visually observed

The shear zone may be more complicated than a simple main shear near the footwall with

gradually decreasing intensity towards the hanging wall. Based on geochemical evidence

there is some reason to believe that the shear zone has either been active to varying degrees

at different levels in the sequence, or that different planes were active at different times.

In addition in a number of holes a basaltic dyke has been intersected which may offset the ore

body however a good 3D understanding of this dyke does not yet exist.

The main ore zone is characterised by the following features

potassic alteration



silicification

small irregular sulphides, mainly pyrite

retrograde amphibolite facies

greenschist facies



Figure 7.3_1 Geology of the Posse Deposit



7.4 Mineralisation

In general, mineralization at Posse is developed along a 50° to 65° striking fault set.

Mineralization tends to be strongest within mylonitic zones that follow more northerly striking

(approximately 30° to 50°) shear strands and dilatant jogs within the larger 50° to 65° striking

major shears that obliquely transect the contact between the Gray Gneiss in the hanging wall

and the Muscovite Schist and Medial and/or Eastern Amphibolites in the foot-wall (Figure

7.3_1). Shearing of the Gray Gneiss has resulted in the formation of a distinct lithologic unit,

the quartz-feldspar-mica schist (Posse Schist) that is characteristic of the Posse ore zone.

This unit has been identified in several other areas including the Posse foot-wall and on strike

extensions of the Posse Ore Zone to the northeast.

Mineralization at Posse has been traced along strike for approximately 1.5 km with an true

thickness ranging between 15 m and 30 m and has been drilled to a vertical depth of 400 m.

The Posse Deposit was reported by predecessors to be confined to two separate northwest-

dipping tabular lens-like lodes, referred to as the North lode and South lode. Both lodes have

been partially mined by open pit (the North lode pit has been backfilled). Mapping of the South

lode has indicated the presence of a 34° to 40° westerly plunge which has been speculated to

be the result of reverse-dextral oblique slip along the main Posse shear zone (Caddey, 1997

and Rosendo, 1998). This plunge may control the overall trend of gold mineralization in the

deposit, and is considered key to understanding and exploring for shear-hosted gold

mineralization elsewhere in the district. An inclined longitudinal section of the Posse Deposit

facing southeast is presented in Figure 8.4_1. The drill intercepts have been plotted showing

grade times thickness of each intersection (gram-metres). The plunge of the deposit is

highlighted by the contours.



Figure 7.4_1

Inclined longitudinal section of the Posse Deposit



8 DEPOSIT TYPES

Several significant mineral deposits occur in the Mara Rosa region including the Posse gold

deposit, the Zacarias gold-silver-barite deposit and the Chapada copper-gold deposit, in

addition to numerous historic prospects and garimpos (Table 8_1).

Table 8_1

Mara RosaProject

Significant deposits in the Mara Rosa region

Deposit Deposit Class References

Posse Au (Eastern Belt) Shear-hosted mesothermal lode-gold Metallica data (Mara Rosa files)

Amarillo website

Zacarias Au-Ag-Ba (Central Belt)

Stratiform syngenetic exhalative or shear related epigenetic high sulphidation?

WMC data (MR files); Poll, 1994. R. Shaw/M. Petersen

Chapada Cu-Au (Eastern Belt)

Volcanogenic exhalative? Wall rock porphyry copper system ?

Kuyumjian, 1991. Richardson, et. al., 1986; 1988.

The Posse Deposit is hosted in a regional thrust that probably acted as one of the primary

dewatering conduits during the Neo-Proterozoic Brasiliano orogeny. The geophysical,

geological and geochemical data available demonstrate that the Posse Deposit occurs within

a 50 km long structural zone with potassium alteration and lower order gold- copper-

molybdenum mineralization. The Posse Deposit has a hanging wall of grey gneiss and the

foot wall of amphibolites, “greenstone”, and it is speculated that the rheological contrast

between the two rock types captured the regional thrust (movement West to East) for a 2 km

segment. It is also possible that the chemical contrast between these acid hanging wall and

basic foot wall may have aided in focusing the mineralizing fluids. Observations in the core

suggest that an earlier potassic event with chalcopyrite, molybdenum, quartz veining, biotite

and K-feldspar was followed by a later auriferous phyllic event with pyrite, Fe-telluride, and

sericite. The gold occurs free as well as associated with the telluride and pyrite.



9 EXPLORATION

No new exploration work on the Posse mineral occurrence was carried out between the

completion of the 2008 drilling and the completion of the 2010 PEA report. The only

exploration work carried out since then is the 2010/2011 drilling program (the 2011 drilling),

which commenced in October 2010 and was completed in March 2011.



10 DRILLING

10.1 Drill Hole Planning

Drill Hole Planning was carried out in two parts:

1. By Amarillo Staff with the focus on the collection of metallurgical and geotechnical data

2. By Mr. Hoogvliet of HCS with the focus on conversion of Inferred blocks from the 2010

PEA Resource Estimate to Indicated or higher.

Initial hole locations were chosen to collect information on metallurgical and geotechnical

data, these locations were then referred to HCS so that additional holes could be planned

which would maximise the conversion of blocks identified as Inferred in the 2010 PEA to

Indicated while at the same time upgrading Inferred to Measured blocks.

The 2010 PEA block model was displayed showing the class of blocks (1=Measured;

2=Inferred; 3=Indicated). Starting in the southwest on section 1, the most obvious locations

were chosen for additional drilling to both convert class 3 to a higher class, and to confirm

‘bottom of the pit’ (the US$1,000 shell from the 2010 PEA report) mineralization. Additional

selection criteria were applied as discussed in the “Report on Independent Site Visit and

Resource Estimate ; Posse Deposit, Mara Rosa; Goiás State, Brazil; dated 30 July 2011

prepared by Hoogvliet and Whitehouse. The results of this planning study are presented

graphically in Figure 10.1_1.

Based on this study, an estimated 82% percent of ore blocks within the US$1000 pit shell

would be classified as Indicated or Measured, and able to be included in ore reserve

estimates and pre feasibility pit design. This compares to 56% of the blocks in the 2010 PEA

study.



Figure 10.1_1

2011 drill hole plan with the US$1,000 shell (brown) at 420RL

10.2 Technical and Support Staff

With the onslaught of the Global Financial Crisis in 2008, the then current team was

dismantled. The only remaining technical staff member from 2008 was Augusto, an

experienced exploration geologist. For the 2011 drilling a new team was assembled,

consisting of Wilson, and Rodrigo. Support staff from the 2008 drilling was re-hired from the

local population, consisting of Jesus and Jose, both very experienced field technicians with a

significant amount of historical exploration knowledge.

10.3 Drill Hole Setup

Holes are marked up by handheld GPS, but picked up by DGPS. The magnetic azimuth is

marked up with three pegs in-line, setup with a Brunton compass. The difference between

magnetic and true north, according to Augusto, is 20o (in the 2008 campaign it was 19o). This

was confirmed on www.magnetic-declination.com which shows the declination for Goiania to

be -20o 10’. So for a hole drilled with a true azimuth of 140o the markup should be 160o

magnetic. A check on the setup of hole (MRP0035 (plan ID 22, one of the holes about to be

collared during the site visit) showed 157o magnetic (137o true) with a Suunto compass. The

drillers manoeuvre the drill rig according to the setup string to the best of their ability. Before

drilling commences, a geologist from Amarillo does a field check of the drill hole setup. A field

inspection was undertaken during the site visit and a comprehensive check reported in the

“Report on Independent Site Visit and Resource Estimate ; Posse Deposit, Mara Rosa; Goiás

State, Brazil; dated 30 July 2011 prepared by Hoogvliet and Whitehouse.



10.4 Drilling Execution

The Drilling contractor was Mata Nativa Comercio e Servicos Ltda, from Nova Lima (MG),

which utilizes I-800 triple tube wireline drill rigs, equivalent to a Longyear LY44. The three rigs

used on the 2011 drill program were several years old, but generally in good condition. Each

rig was operated by a driller with four assistants. The rigs were all of the same model, and in

similar condition. Three rigs were in operation during the site visit. The drillers as well as the

crew appeared experienced enough to do their job in a competent way.

10.4.1 Down Hole Surveys

Down hole surveys are collected with the Maxibor system, an instrument often used in highly

magnetic rocks. It is a down hole optical survey system which measures changes in vertical

dip and lateral variations. It is controlled from a palm top on the surface, which runs the Reflex

software. It requires input of E, N, RL and Azimuth of the top of the hole (the collar). This

information is downloaded to the survey tool prior to lowering the instrument. It can read in

intervals of 1.5 m or 3 m, and read from top to bottom, bottom to top, or both. It does not

measure azimuth as it does not have a compass.

Upon completion of the survey, the data is downloaded onto the palmtop and sent to a

Maxibor office for conversion of the data to E, N, RL for each 1.5 m or 3 m reading. This file

then gets send to the client. The survey takes about 15 minutes for every 100 m. If preliminary

data are used for E, N, RL and/or azimuth, the downhole data needs to be re-calculated when

final location information becomes available.

The Maxibor tool used at Posse was owned by Mata Nativa, the drilling company doing the

current drilling.

10.4.2 Driller’s Field Records

The field recording by the driller includes: From (m), To (m), Drilled (m), Core Length (m),

Length of Rod (m). Initially this information is recorded in a field book, which at the end of the

shift gets transcribed into the official log, which forms the basis for approval and payment.

The drill data records from the driller are not separately captured in a file.

10.4.3 Core mark up (Field)

As the drill core is extracted from the core tube, it is laid out in a 3 m long V-shaped tray, from

which the data above are collected. The core is cleaned with a cloth and water, the relevant

drill data is recorded, the core is transferred to core boxes with 4 or 5 bays (depending on

core size), from the top left to the bottom right. The individual runs are clearly marked by

wooden core markers, with the depth at the end of the run punched into aluminium tags,

which in turn are nailed to wooden blocks. For each run a spear is lowered to mark the base

of the core. Upon retrieval of the core, the imprint of the spear is highlighted and the pieces of

core are re-assembled into their original relative position, and a reference line is marked on



the piece of core with the spear imprint. Each core box is clearly marked with the hole number

and other relevant details.

The core was transported to the core shed near Mara Rosa on a daily basis by the drill

contractor.

10.4.4 Collar Preservation

After the hole is completed and the rig moved to another location, a piece of PVC pipe is

cemented in at the location of the hole, with the hole number, start and end date and the

depth clearly marked. All rubbish is removed; the site is levelled, and rehabilitated, or left to

rehabilitate itself. Most collars have a cap on the PVC pipe.

10.4.5 Core Shed Procedures

The individual runs of core are re-assembled in a V-shaped tray, with the individual pieces

matched to each other, and the bottom of the core (based on the position of the spear mark)

is marked along the core for its entire length. This procedure ensures that the core is cut

without bias towards veins and structures. It also allows for collection of oriented core data.

Magnetic susceptibility data is collected with a frequency of one spot reading per metre. It

should be noted however that a search for historic magnetic susceptibility measurements has

only revealed data for holes F001 to F009.

The geotechnical data was collected and before cutting, the dry core is photographed.

The depth of the core is marked in 1 m or 2.5 m intervals on the side of the bay, and the

sample number is punched into an aluminium tag, which is nailed underneath the relevant

depth marking.

10.5 Densities

Density measurements were collected as part of the 2011 drilling program however at the

date of this report the most recent measurements had not been validated so the results have

not been included in this Technical Report. It is however noted that initial results suggest a

slightly higher SG than was used in the resource estimation reported in Section 14 of this

report. Amarillo are also preparing a set 200 of samples which will be used for density

determination by a recognised independent laboratory. It would be expected that the results of

this work will be used for density estimates in the future.

Historically there have been a number of density measurements used based on the sets of

data listed below :

Set 1: The 2008 Amarillo data over three MRP holes (10-12)

Set 2: The 2007 CCIC samples



Set 3: the WMC data from the mid 1980’s

Set 4: The 2006 Zonge data

The data in each set of measurements is summarized in Table 10.5_1.

Table 10.5_1

Mara Rosa Project

Historic SG measures

Zonge CCIC Amarillo WMC

Mean (all) 2.77 2.78 2.56 2.85

Median (all) 2.72 2.77 2.55 2.82

Mean (>0.4 g/t) 2.71 NA 2.57 2.85

The 2010 report discussed the validity of each set of data and the authors concluded taking

into account the uncertainties on the actual determinations, the quality of the Zonge data as

well as the theoretical determinations, that a density of 2.73 across the board was the most

appropriate at this time. This report has adopted the same value for density.

10.6 Sampling Method and Approach

Prior to cutting the core a short log is created to assist sampling. The core is then cut on site

using the line marked on the core as a guide. The actual cut is on one side of the line,

resulting in one half being exactly half, whilst the other half (which is used for sampling) is less

than half, the difference being the width of the blade. This cut is achieved by having a slightly

off-centre sledge, or tray, in which the core rests during cutting. This procedure is on the

advice of AVB (geotechnical consultants).

The core is sampled at 2.5 m intervals in waste areas (away from the main zone) and 1 m

intervals in the main zone.

10.7 Geological Data Collection

Detailed logging is conducted on site. However the logging at the time of the site visit was

lagging well behind drilling. The logged data is entered into an Excel spreadsheet and prior to

final acceptance the logging is checked and the holes are manually plotted on sections.

Logging includes:

Lithology: 16 codes

Alteration minerals

Sulphides: pyrite, pyrrhotite, chalcopyrite

Minerals: magnetite, garnet, chlorite



Long hand description

10.8 Opinion

Coffey Mining is of the opinion that the drilling program, and all procedures, have been

conducted in accordance with industry best practise, and that the results are both relevant

and suitable for purpose of mineral resource estimation. The combination of lithological log

and survey data ensures that only true widths are used in the estimations.



11 SAMPLE PREPARATION, ANALYSES AND SECURITY

Upon completion of the cutting, samples are collected in plastic bags. Each sample contains

two tickets from a pre-printed sample number book. In the book, the details for each sample

are recorded, including hole # and from-to. Regular samples are ‘half’ core, whilst the half

core is recut to produce 2 quarter sections of core for duplicate samples. Prior to the 2010 –

2011 drilling program no duplicate samples were collected however on the recommendation

of HCS and AEFS in the 2010 Report a decision was made to generate duplicates as part of

the QA/QC sample stream. The results of this work and the other QA/QC sampling are

discussed in section 12.2.1 below.

Samples are despatched to the laboratory in lots of 150.

The plastic bags containing each sample are packed into canvas bags with 4-5 samples in

each bag for transport to the laboratory. The canvas bags are sealed and marked with the

requisition number, address, sender’s details, and marked 1/28, 2/28 etc. One copy of the

requisition is retained and stored in the company archives and another goes to the driver

taking the samples to the laboratory. A copy of the requisition is also sent to ACME, the

laboratory carrying out the assay work, and Amarillo staff by email.

Initial sample preparation is carried out in Goiania by ACME labs. The procedure at ACME

includes the following steps:

Sorting and checking against the requisition sheet

Drying at 60o C

A granite wash is used to scour the equipment before the client’s first sample is crushed

Crushing of the samples to 80% passing 10# (2 mm)

Samples are homogenised, and riffle split to 250 g subsample

A granite wash is used to scour the equipment before the client’s first sample is

pulverised

Client’s samples are pulverised to 85% passing 200# (75 µm)

Equipment is cleaning by brush and pressurised air

A granite wash is used to scour the equipment after high grade samples, between

changes in rock colour, and at the end of each file.

The ACME sample preparation facility in Goiania was visited by Kevin Kivi, an experienced

exploration geologist from Canada. The procedures were reported to be of a high standard

and the results of QAQC sampling support this even though the ACME sample preparation

facility in Goiania is not accredited to ISO 9001.



Approximately 100 g of the final pulverised samples, standards and blanks are sent to the

assay laboratory, with a pulverised duplicate retained at ACME. Assaying is carried out in

Canada by ACME’s central assay laboratory in Vancouver which was accredited under ISO

9001 in November 1996. The registration has been maintained in good standing since then.

The pulp duplicates, are stored by ACME in Goiania for 3 months before they are returned to

Mara Rosa. A flow chart of the sample preparation is presented in Figure 11_1.

Gold assay was by a 30 g fire assay with ICP finish. The laboratory has a 10 g/t Au upper

detection limit with over range samples being re-assayed with a gravimetric finish if the initial

assay result was greater than 10 g/t Au.



Figure 11_1

ACME sample preparation flow chart

It is the opinion of Hugo Hoogvliet of Hoogvliet Contract Services that the 2010 – 2011 drilling

program and all its associated procedures is of a sufficiently high quality that the data

obtained from the program can be considered to be reliable and included in this 43-101

technical report.



12 DATA VERIFICATION

12.1 Data entry

The logging spreadsheets were used to generate a set of csv files prior to import into

Micromine. The imported data showed a number of errors, largely related to missing or

overlapping From and To’s, and inconsistencies between bottom of the hole in the collar data

and the From and To’s. These errors were corrected prior to modelling work being

undertaken.

12.2 Amarillo QAQC

QAQC samples are an important part of the data verification process and should be routinely

used and analyses as part of any drilling program. At Posse QAQC samples generally

account for more than 10% of all samples. Initially QC followed 2008 the procedure, which

included one standard (STD), one blank but no duplicates every 20 samples. This procedure

was modified from MRP0018 to provide for the inclusion of duplicate samples as

recommended in the 2010 report. QAQC samples, consisting of a standard, a blank or a

sample duplicate, are now inserted every 25 samples on average. This allows a degree of

flexibility in placing the QAQC samples so that they are used to the best advantage.

Standards and blanks have been supplied by Rocklabs from Auckland, New Zealand.

Standards are shipped in jars of 2.5 kg. A small sample is scooped out, and weighed to get

100 g and put in a plastic bag, apparently the same type as those used in the sample

preparation lab in Goiania.

Blanks come in 50 g packets; two are combined into one to make up one 100 g blank sample.

Certificates for all three standards have been sighted:

SK43 has an average grade of 4.086 g/t Au with 0.036 g/t variation for 95% accuracy

SJ39 has an average grade of 2.641 g/t Au with 0.033 g/t variation for 95% accuracy

SH35 has an average grade of 1.323 g/t Au with 0.017 g/t variation for 95% accuracy

12.2.1 QAQC Results

As part of the data collection process regular standards and blanks were inserted into the

sample stream. Plotting the expected versus the actual value of a standard or blank over time

will indicate any major deviations from the expected value. When this happens the cause can

be investigated and if necessary the batches of assays associated with the out of control

standards can be submitted for re-assay. Throughout the 2010 - 2011 drilling program three

Standards were used. The standards used were all supplied by Rocklabs of Auckland, New

Zealand and were coded SH35, SJ39 and SK43. The certified value of the Standards is set

out in Table 12.2.1_1.



Table 12.2.1_1

Mara Rosa Project

Certified values of standards

Standard Certified Value

SH35 1.323 g/t

SJ39 2.641 g/t

SK43 4.086 g/t

Run charts covering the period of the drill program show samples outside the expected +/-

10% boundary. In the case of Standard SH35 it would appear as though two blank samples

have been miscoded as SH35. The chart of SJ39 shows one outlier, this has a value of

around 4 g/t and is probably actually a sample of standard SK43 that has been miscoded. The

chart of SK43 has all but one sample within the control bands, the one point which is out is

only just out of bounds and does not suggest any serious issues with the QAQC data.

In addition to the above standards, blanks and sample duplicates were also used for QAQC

sampling. The blanks span the period of the 2008 drilling and the 2010 - 2011 drilling and

show a major change with the start of the 2010 – 2011 drilling program. The 2010 – 2011

drilling program shows a much lower incidence of blanks returning higher than expected

assay results (i.e. Assay results which are more than the assay detection limit) and indicates

a major improvement in the quality of the sampling and recording of data associated with the

latest drilling program.

The duplicate samples which were taken for holes MRP0018 onwards would be expected to

show an extremely high correlation with the original sample values. A scatter plot of the

original samples versus the duplicates is the standard way of analysing such data and in this

case indicates a correlation of better than 98% with no obvious outliers. This is a very good

result and indicates that the data was well recorded and that the assay results are repeatable.

The repeatability of the assay results not only indicated that the laboratory is doing its job

correctly but that there are no large issues with nugget gold where coarse grains of gold lead

to a lack of repeatability in the assay results.

The results of the sample QAQC program associated with the samples from the 2010 – 2011

drilling indicate a generally high quality of work and do not show any cause for concern.

12.2.2 Due Diligence QAQC

As part of the site visit the Mr Hoogvliet collected 32 samples consisting of ¼ core duplicates,

pulp duplicates obtained from ACME in Goiania, as well as 2 blanks, and two samples each of

the three standards. The results show excellent correlation between the original assay results

and the repeat assay results.



12.3 Drillhole coordinates

One of the recommendations in the 2010 report was that all new drillhole be accurately

surveyed and that where possible historic collars should also be located in order to compare

the historical coordinates with modern accurate coordinate determinations. In order to

facilitate this Amarillo have acquired a high quality differential global positioning system

(DGPS) receiver. This instrument can measure horizontal coordinates to a precision of +/- 0.1

m or better. Vertical coordinates are less well defined but are within + / - .5 m. All drillhole

collars from the 2010 - 2011 drilling campaign were located with the DGPS together with other

drillhole collars which could be physically located in the field. Final measurements were made

using the WGS 84 datum and coordinates were recorded in terms of a UTM projection, Zone

22S (WGS84 coordinates), these coordinates are effectively the same as coordinates in the

Zone 22S projection of the SIRGAS2000 datum. From 2014 the SIRGAS2000 datum will be

the only legal datum used in Brazil and the SAD69 datum is being phased out.

Coordinates of all holes in the 2010 - 2011 drill program were also recorded in SAD69

coordinates. Since all the drill data is located in a small (geographic) area a plan grid

transformation based on the average difference in East and North of all holes located in both

the WGS84 and SAD69 datum's provided a simple transformation that could be applied to

convert the locations of holes and other data where the coordinates were only known in the

SAD69 to the WGS84 datum as shown in Table 12.3_1.

Table 12.3_1

Mara RosaProject

Coordinate conversion SAD69 to WGS84

SAD 69 to WGS84

East -43.57 m

North -43.04 m

12.4 Topographic survey data

The 2010 report used a topographic surface derived from digitized topographic maps

referenced to the old Corrego Alegre datum. The maps were based on 1985 aerial

photography with some ground-truthing. Comparison of the digitised data with drillhole collars

etc. indicated a good fit to the pre-mined surface. With the arrival of a high quality DGPS on

site at Posse in early 2011 the opportunity is being taken to update the topographic surface

model. This work has involved surveying all open areas surrounding Posse with a DGPS and

substituting valid X, Y and Z data points obtained from the DGPS for the points in the same



areas on the digitized topographic surface. Data used for elevation observations is limited to

DGPS observations with a PDOP <31.

The revised surface topography provides a good fit to historic collar elevation values, it is

expected that further work will be carried out to verify the surface topography as the project

progresses into Prefeasibility and it is hoped, Feasibility.

12.5 Downhole Survey

As part of the data review conducted as part of the validation process it was noted that the

holes in the 2010 - 2011 drilling program from MRP0020 - MRP 0045 all showed a noticeable

deviation, in all but one of the holes this was to the south (a clockwise turn when looking down

the hole). Drill hole MRP0030 was the only one which turned anticlockwise (Figure 12.5_1).

Figure 12.5_1

MRP series holes

The reason for this deviation in the holes is not clear and Amarillo would be advised to check

the actual trajectory of the holes with a multi-shot survey if the holes can be accessed. It

1 PDOP refers to Position Dilution of Precision as is a standard measure of the quality of GPS data. A lower PDOP indicates a better satellite geometry and hence reliability of data and accuracy of coordinate observations.



should be noted however that when the mineralisation wireframes were generated particular

attention was paid to these holes to check if they caused any disruption or irregularity in the

wireframes. No such effects were observed and this would suggest that the deviations

observed are real. The deviation of the holes has meant that a number of the planned holes

did not end up in the location planned. This was examined and taken into account when

blocks in the mineralisation model were coded for resource classification.

12.6 Improvement of Drilling Programs

The 2010 – 2011 drilling campaign meet the objective of providing additional information in

relation to the mineralization at Posse. There are however a number of issues related to the

setup and management of the drill program which if addressed will improve the quality of data

gathered. Where relevant the suggestions outlined should be adopted for all future drilling at

Posse.

12.6.1 Drill Rig Setup

Mr Hoogvliet noticed that during the field check of the rig setup before drilling commenced,

the mast was in a vertical position, and the skids of the rig were perpendicular to the drill

direction. This removes the two most reliable reference points to properly setup the rig.

Lowering the mast to a horizontal position while sighting would allow the geologist to better

gauge how well the rig is aligned with the sighting string. Similarly, reading the dip on the

Brunton compass with its relatively small dials, is not ideal. Use of 50 cm clinometers would

make the setup more accurate, and easier.

12.6.2 Down Hole Surveys

The point made under drill rig setup carries through to down hole surveys. Because the down

hole surveys were done with an optical instrument (Maxibor) which requires azimuth as input,

the accurate setup of the rig is therefore more important than if a magnetic down hole tool

were to be used. Any inaccuracy introduced during the setup of the rig carries through to the

down hole surveys.

In the 2010 – 2011 drilling program the setup at the collar is dependent on a handheld

compass for both azimuth and dip. The inaccuracies inherent in this setup will therefore limit

the accuracy of the surveys obtained from the Maxibor survey which is designed to be used in

a situation where the rock being drilled is highly magnetic. This is not the case at Posse and it

is therefore recommended that subsequent to drilling, standard downhole survey data is

collected with an Eastman, Tropari or equivalent instrument where possible.

12.6.3 Data base

Because the Maxibor is provided with true collar and down dip information for its survey, there

is no survey of the magnetic dip direction. However, magnetic ‘data’ is recorded in the data

base as a calculated number. In data from the 2010-2011 drilling program, the 20o difference

between magnetic and true is subtracted for some holes in the data base, and added to other



holes in the data base. Even though the determination of the magnetic azimuth is in this case

mostly academic, part of the magnetic azimuth ‘data’ is incorrect. If as recommended above

future downhole survey work use magnetic north based measures of azimuth and dip then it

is essential that this is calculated correctly.

The website Magnetic declination.com reports the declination as -20o27” at Mara Rosa, so to

go from a true azimuth to a magnetic azimuth requires the declination to be subtracted,

similarly to go from magnetic to true azimuth the declination must be added. In the data

provided Holes MRP0015 – MRP0020 had magnetic azimuth correctly calculated; the others

were incorrect.

12.7 Drill Program Assessment

It is the opinion of Mr Hoogvliet that the 2010-2011 drilling program was executed to adequate

or higher standards, including the associated sampling, surveying, assaying and quality

control procedures, and is of sufficient quality to be included in a 43-101 compliant report.



13 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 Introduction

Coffey Mining were initially requested to undertake a review of the preliminary metallurgical

testwork results and data from the Mara Rosa project, prior to being appointed to develop and

manage a suitable metallurgical testwork programme for the project to a pre-feasibility level.

The full results of this review and testwork program are included as Appendix A of Smith and

Witt (2011).

Prior to the Coffey Mining review, Amarillo had commissioned a preliminary metallurgical

testwork study with the aim of establishing likely gold recovery and flowsheet options for

processing mid to low grade Posse mineralisation. Previous higher grade (~3.5 g/t Au) Posse

mineralisation processed by WMC yielded recoveries of approximately 84% using standard,

aerated CIL techniques.

The preliminary metallurgical testwork aimed to establish indicative gold recoveries that might

be obtained using similar processing techniques to those previously used, but at a significantly

lower resource grade of approximately 1.5 g/t Au.

The Coffey Mining review noted that the gold mineralisation was associated with tellurides

and sulphides, and that previous testwork programmes had failed to recognise the

fundamental requirements for recovering gold from such minerals, hence the testwork

recoveries tended to be inconsistent and less than optimal.

The recovery of gold from refractory sulphides is dependent on the association of the gold to

the sulphide. If the gold is simply occluded (or locked within) the sulphide, then the recovery

of such gold can be as simple as ensuring that the sulphides are ground fine enough to

liberate the gold so that it can be dissolved in cyanide. If the gold is part of the atomic

structure of the sulphide species, then it can only be recovered if the sulphide structure is

broken down (generally oxidised) releasing the gold so that it can then be dissolved in

cyanide.

However, for gold tellurides, it is imperative that the telluride be oxidised first, else the gold will

remain insoluble in cyanide solutions. In a normal agitated leach tank, this may take up to 48

hours or more at an elevated pH of around 12. Only after the telluride has been oxidised can

the gold be dissolved in cyanide.

13.2 Background

A number of metallurgical testwork investigations have been conducted on the Mara Rosa

mineralisation, primarily by WMC and Barrick.

Prior to the commissioning of the more recent testwork programmes, a number of preliminary

cyanide leaching tests were completed on Posse core pulp samples at the metallurgical

laboratory and pilot plant run by the Goias State mining entity Funmineral in the city of Goiania.



A number of mid to low grade samples were prepared from samples taken from 11 drillholes

drilled within the Posse Deposit with grades of 0.60, 0.76, 1.09 and 1.39 g/t Au. The drill core

had been previously pulped by Acme Labs in Goiania so that the samples had a P90 of

approximately 75 µm. The results of the Funmineral tests indicated that an average gold

recovery of 83% could be obtained from standard cyanidation bottle roll tests. It was

suspected that the higher than expected gold recoveries might have been due to the natural

oxidation of the sulphide materials during storage.

Amarillo commissioned Desenvolvimento de Processo Ltda to complete a new testwork

programme in December 2009 to determine the likely metallurgical performance and gold

recovery of the Posse sulphide mineralisation at a lower cutoff grade of 0.5 g/t Au and a

predicted run of mine gold grade of 1.48 g/t Au. The testwork incorporated gravity and

flotation concentration followed by cyanide leaching, including pre-oxidation treatment.

A total of 212 kg of drill core intercept samples from the Posse mineralisation zone were

selected and from these, six composite samples representing the main lithological domains

and likely plant feeds were prepared.

Preliminary results from the cyanidation testwork yielded similar results to previous testwork in

terms of lime demand, cyanide consumption and recovery. Under standard CIL conditions a

consistent solid residue of between 0.40 g/t – 0.60 g/t Au was observed, which equates to a

recovery of 60%-75% at a head grade of approximately 1.48 g/t Au.

During the subsequent kinetic and pre-oxidation tests however, several anomalous results

from one particular composite (Sample F) resulted in the suspension of the testwork. It is

unclear whether the erroneous grades reported were the result of representivity issues

relating to the splitting of the composite sample, or possible analytical error at the laboratory.

In order to further investigate the issue, the metallurgical laboratory subsequently re-

homogenized the anomalous composite sample and re-assayed the material. The results

returned were as expected, however, due to questions regarding the sample’s representivity,

the testwork programme was halted.

Flotation testwork was carried out and achieved gold recoveries from 50% to 79% to a

sulphide concentrate whilst preliminary Knelson centrifugal gravity concentration tests

recovered less than 30% of the gold at a centrifugal force of 60 g.

At this point, Amarillo provided Coffey Mining with summary reports of the historical testwork

results for review. The majority of the testwork results related to mineralogical analysis,

cyanide leaching and flotation recovery performance.

Mineralogical analysis indicated that the Posse mineralisation consisted predominately of pyrite

with minor chalcopyrite, pyrrhotite and galena. Gold was present as both free grains and also

associated with tellurides (predominately iron tellurides FeTe2). Tellurides occurred as isolated

grains disseminated in the gangue or included in pyrite crystals. The main telluride minerals



were calaverite (AuTe2), forming complex intergrowths with stibnite and native gold, and

frohbergite. Sylvanite (AuAgTe4) was locally observed, associated with calaverite. The gold

grain size was generally fine, ranging from 1 µm to 10 µm and occurred in various locations,

including within silicates; composite with iron telluride; and/or at silicate grain boundaries.

Cyanide leach testing indicated that the Posse mineralisation contained a refractory component

that was not readily cyanidable under normal conditions. Solid residues in the range of

0.4 g/t Au – 0.6 g/t Au were obtained in the majority of the Posse sulphide (deep) samples. This

equated to a recovery of 60% - 75% at a 1.5 g/t Au head grade. Some samples of Posse oxide,

or shallow mineralisation, yielded significantly better recoveries in excess of 95% under

standard leach conditions.

Kinetic leach tests indicated that the sulphide mineralisation contained a fast leaching, readily

cyanidable portion, followed by an extremely slow leaching, semi-refractory portion. This was

consistent with the mineralogical observation of both fine, free gold and gold/telluride/pyrite

phases.

Diagnostic testwork conducted on the leach residues indicated that the majority of the non-

leached gold was associated with the pyrite phase and to a lesser extent pyrrhotite and

silicates, however, in the presence of tellurides, such diagnostic leach testing can often be

flawed. It has been estimated that on average the mineralisation contained approximately

65% readily cyanidable gold and 35% refractory gold.

The diagnostic tests performed were unable to differentiate between the refractory sulphide

and/or telluride components of the mineralisation. It was unclear if the refractory portion of

gold was fine gold locked within pyrite, gold telluride species, gold/telluride locked in pyrite or

a combination of all three.

A number of standard methods were investigated in order to increase gold recovery, including

increasing cyanide concentration, increasing residence time, reducing the grind size, the

addition of lead nitrate and pre-oxidation. Of these only the intensive pre-oxidation had any

significant effect in increasing the gold recovery, yielding >80% @1.5 g/t Au head grade,

although results were inconsistent.

Preliminary flotation testwork by WMC indicated that >80% of the gold and >90% of the

sulphur could be readily concentrated by flotation under typical conditions. The testwork

performed involved simple single stage rougher tests. No locked-cycle, cleaning or reagent

optimisations were performed.

WMC also operated a high grade CIL and heap leach operation processing the Posse

mineralisation in the early 1990’s, with approximately 80,000 oz Au being mined from the

deposit at an average grade of 3.5 g/t Au. Summary historical production records provided by

Amarillo indicate that the standard CIL circuit returned an overall gold recovery of 84% at this

higher head grade. This would appear to be consistent with the testwork above and indicates

that the solid residues were in the order of 0.50 g/t Au to 0.60 g/t Au.



Coffey recommended that Amarillo undertake further testwork in a staged manner, with

decisions for each subsequent stage based on the previous stage outcomes. This would

enable design criteria and flow sheet options suitable for a pre-feasibility level of study to be

established.

Although not an imperative to success, Coffey Mining recommended that the testwork be

carried out at a commercial laboratory in Perth, Australia, where previous telluride testwork

has been carried out with successful outcomes.

The presence of gold telluride is less common in operating gold projects possibly due to its

refractory nature and the poor understanding of how to successfully process material that

contain such mineralisation. In simple terms, gold tellurides are insoluble in cyanide unless

they have first been oxidised prior to cyanidation. Some oxidation reactions for gold telluride

are shown below.

AuTe2 + 2O2 + 2H2O ↔ Au + 2H2TeO3 (aqueous)

AuTe2 + 2O2 ↔ Au + 2TeO2 (gaseous – roasting)

In a normal agitated leach tank, the aqueous process may take 48 hours or more, even at

elevated pH (>12), which increases the oxidation rate. Gold telluride that is not oxidised does

not leach in cyanide and can therefore be the reason for reduced gold recovery in certain

samples.

Based on the information provided by Coffey Mining, Amarillo instructed Desenvolvimento de

Processo Ltda to complete some further metallurgical testwork with a focus on finer grinding

of the samples and pre-oxidation of the gold telluride prior to the cyanidation step. A P80 grind

of 38 µm was targeted for each of the samples and then calcium hypochlorite at 5%w/v was

used to pre-oxidise the samples over a 4 hour period prior to leaching. This methodology of

finer grind, pre-oxidation and leaching generally achieved gold recoveries in excess of 90%

and up to 98%.

Although the outcomes of the additional testwork were greatly improved in terms of gold

recovery, the use of calcium hypochlorite (particularly at such concentrations) as a per-

oxidant is impractical in a full scale plant. However, the important component of these

testwork results was that they verified Coffey Mining’s opinion that the gold tellurides needed

to be oxidised if they were to be soluble in cyanide.

As an aside, calcium hypochlorite is a very strong oxidant, but is also similar in price to

sodium cyanide, which is used at around 0.030% to 0.050% w/v in most gold plants. Hence

the use of calcium hypochlorite at levels of 5% w/v in the gold recovery process is not realistic

in terms of process flow sheet development.



There are relatively few gold operations processing gold telluride, and as a consequence,

there is relatively limited knowledge of the testwork requirements in many laboratories (and

consultancies) worldwide.

An important component of the Mara Rosa testwork samples would likely be a comprehensive

understanding of the mineralogy and nature of the cyanidation leach residues. The use of

correct procedures and competent personnel would be vital in this area.

In addition to this, there have been several processes that have been developed and refined

in more recent years that are potentially capable of providing an economic means for

treatment of difficult refractory gold mineralisation, for example, the Albion process. For these

reasons, it was strongly recommended that certain components of the testwork be carried out

in specific laboratories rather than generic ones.

Amarillo decided to send approximately 316 kg of samples representing the foot wall, main

zone and hanging wall to Perth Australia, where laboratories were reasonably familiar with the

requirements for processing difficult refractory gold ore samples.

13.3 Metallurgical Domaining

Discussions were held between Amarillo and Coffey Mining regarding the nature of the

deposit from a mineralogical and metallurgical perspective. Based on the known geology and

the preliminary mineralogical testwork that had been previously undertaken, it was decided

that the most suitable mineralogical domaining for the proposed testwork programme would

likely be based on the following:

Foot wall domain FW ~10% of the deposit

Main domain Main ~60% of the deposit

Hanging wall domain HW ~30% of the deposit

With subsequent changes in the resource base these proportions are now:

Foot wall domain FW ~3% of the deposit

Main domain Main ~82% of the deposit

Hanging wall domain HW ~15% of the deposit

From a geological perspective, the deposit displayed reasonably good homogeneity and so

the domaining as described above may not be required. Lithologically, the samples had been

categorised by position (FW, Main, HW) and by colour (grey, pink, black).

The Coffey Mining review had identified that the most likely problematic components to the

Posse Deposit would be sulphide and / or telluride gold species. It was decided that an

approach towards the non-sulphide mineralogy would be applied, that is, do the samples

having different colours that represent the FW, Main and HW display different mineralogies on



a macro scale. For example, is the mineralogical make-up of the MAIN samples similar when

a black and pink sample are analysed. Six samples were selected for mineralogical analysis

by Roger Townsend and Associates.

13.4 Sample Selection and Head Grade Analysis

Amarillo dispatched approximately 316 kg of samples from 12 different diamond drillholes

located throughout the deposit. Full multi-element analysis for Main, Hanging Wall and

Footwall composites were undertaken and reported.

13.5 Testwork Programme

The metallurgical testwork programme for the Mara Rosa composite samples was completed

using a staged approach. This was primarily due to the historical testwork results, which had

on occasions produced inconsistent results, largely due to a lack of understanding of gold

telluride chemistry.

Although all three mineralogical domains were tested for their metallurgical characteristics,

there was an initial focus on the main and hanging wall samples, as they represented the

major component of the deposit and the mineralogical analysis did not identify any significant

differences that would likely affect the metallurgical behaviour of the three nominated

mineralogical domains.

The staged approach to the testwork programmes allowed for ongoing decisions to be made

with respect to the necessity of all of the proposed testwork.

In summary, the proposed testwork programme was to include the following:

Sample Preparation:

Select individual samples for mineralogical analysis;

Main and hanging wall composite samples for testwork;

Foot wall composite sample for testwork.

Mineralogy:

Minerals present;

Mineral associations;

Liberation size for relevant minerals;

Gold and telluride deportment (if feasible).

Head grade determination of all three mineralogical domains.

Grind Determination:

At a P80 of 75 µm and 45 µm.



Lime Demand:

Lime required to achieve pHs of 9.0, 9.5, 10.0, 10.5, 11.0, 11.5 and 12.0 at a P80 of

45 µm.

Pre-oxidation and cyanidation time leach testwork:

24 h and 72 h pre-oxidation with cyanidation time leach testwork at 2, 4, 8, 18, 24

and 48 h at 40% w/w solids, pH 12, DO >20 ppm;

Assay leach residue.

Leach residue diagnostic analysis:

If Au in residue >0.10 ppm Au, then complete.

Flotation testwork:

If Au in residue >0.10 ppm Au, then complete if justified.

Optimisation testwork on pre-oxidation and cyanidation time leach (pending results).

Physical testwork:

Ai, Wi(ball), Wi(rod).

Geochemistry on tailings sample.

The details of the Ammtec testwork programmes for the Main, Hanging Wall and Footwall

composite samples are provided in Appendix A of Smith and Witt (2011).

13.6 Comminution Testwork Results and Interpretation

The comminution testwork programme identified the abrasion and grinding characteristics of

the main and hanging wall composite samples. The foot wall composite sample was not

tested due to its low percentage contribution to the deposit.

The samples displayed a medium to high level of abrasiveness and were in the soft to

medium range of competency with respect to grinding.

Table 13.6_1 shows the key design criteria for comminution based on the testwork results.

Table 13.6_1

Mara Rosa Project

Key Design Criteria for Comminution Testwork

Test Unit Main

Composite Hanging Wall

Composite

Ai Abrasion Index 0.3410 0.3426

Wi(rod) Bond rod mill work index kWh/t 13.4 13.1

Wi(rod) Bond rod mill work index kWh/t 12.9 13.0



Diamond drill core samples with a minimum diameter of 50 mm were not available and hence

unconfined compressive strength (UCS) and crushing work index (CWi) testwork were not

completed.

The options of primary crushing / SAG milling versus three stage crushing / ball milling were

considered with respect to the comminution testwork for pre-feasibility level.

Given the metallurgical need for a relatively fine grind and the consequent power requirement,

as well as the lack of clayey material in the deposit, it was proposed that three stage crushing

and ball milling would provide an improved economic outcome, particularly with respect to

grinding power costs. For this reason, any additional physical testwork that would be required

to progress the study to a more detailed level could be completed at that point in time.

Ideally, the UCS would have been determined during this testwork programme, however, the

assumption for the design criteria is that it exceeds 100MPa and hence a crusher capable of

reducing competent rock in the higher ranges of UCS has been proposed.

If SAG milling is to be considered at more detailed levels of study, then JKTech style testwork

would be required.

The abrasion index for the main and hanging wall composite samples were 0.3410 and

0.3426 respectively. This places the samples in the medium to high range for abrasive

properties. For reference purposes, 0.10 would be considered to be in the lower range, 0.25

in the medium range and 0.40 (equivalent to quartz) in the higher range.

The detailed results of the abrasion test are included in Appendix A of Smith and Witt (2011).

The Bond rod mill work index for the main and hanging wall composite samples were

13.4 kWh/t and 13.1 kWh/t respectively.

The Bond ball mill work index for the main and hanging wall composite samples were

12.9 kWh/t and 13.0 kWh/t respectively.

These results place the samples in the softer to medium range of competency for grinding.

Although the option of SAG milling has not been selected in the process flowsheet, the ratio of

the rod:ball mill work indices would suggest that the samples are amenable to SAG milling.

This is based on the fact that the rod:ball mill work index ratios are less than 1.25:1. This

indicative outcome does not preclude the need for further detailed grindability testwork,

should SAG milling be considered as part of the flowsheet in the future.

The detailed results of the Bond rod and ball mill tests are also included in Appendix A of

Smith and Witt (2011).



13.7 Metallurgical Testwork Results and Interpretation

The metallurgical testwork results for the main, hanging wall and foot wall composite samples

indicated that a leach residue of 0.10 g/t Au or less could be readily achieved via a process

flowsheet that included grinding to a P80 of 45 µm, pre-oxidation over a period of 12 h at a pH

of 12 and leaching at conventional cyanide concentrations for a period of 24 h.

Although percentage gold recoveries are displayed in all of the testwork results, these are

invariably grade dependent in most gold deposits, and hence the real focus of the testwork

was to gain an understanding of the nature of the non-recovered gold in the cyanidation leach

residue.

If cyanidation leach residues of 0.10g/t Au or less could not be achieved, then detailed

diagnostic testwork on the leach residues was planned. A leach residue of 0.10 g/t Au or less

was achieved for all three mineralogical domains.

In the case of the main and hanging wall domains, lower leach residues of 0.06 g/t Au were

achieved when the pre-oxidation and cyanidation leach times were extended to 72 h and 48 h

respectively. The foot wall composite sample did not have extended pre-oxidation and

cyanidation leach time testwork performed due to the good results achieved at lesser times

and the lower contribution of this domain to the overall deposit.

The respective head grades for the main, hanging wall and footwall composite samples were

1.39 g/t Au, 1.43 g/t Au and 0.76 g/t Au respectively, whilst the respective leach residue

grades with 12 h pre-oxidation and 24 h cyanidation leaching were 0.10 g/t Au, 0.09 g/t Au

and 0.10 g/t Au, for equivalent gold recoveries of approximately 93% at a head grade of

1.47 g/t Au.

The testwork indicated that under typical CIL conditions (24 h @ pH 10) without pre-oxidation,

leach residues of 0.26 g/t Au and 0.53 g/t Au were achieved, equating to gold recoveries of

64% and 82% for the main and hanging wall composite samples respectively.

This indicates that the effect of the gold tellurides in the different samples can differ somewhat

and if not sufficiently oxidised, then the results may well be erratic and inconsistent, as was

the case in much of the earlier testwork on the Posse Deposit.

The parameter that has the most significant effect is the oxidation of the gold tellurides. There is

minimal literature regarding the nature of gold telluride oxidation (as compared to other gold

mineralisation) and hence the understanding of this subject has remained more of a practical

one learnt in the few operations that treat such materials worldwide, including the Fimiston mine

in Kalgoorlie operated by Kalgoorlie Consolidated Gold Mines (KCGM) that Coffey Mining is

familiar with. More recent developments in refractory gold processing (Albion Process) has

indicated that gold tellurides can be rapidly oxidised (in 2 h to 4 h) at temperatures in excess of

60 °C and at a pH of 9.0 to 9.5. However, without a sufficient concentration of oxidising



sulphides (for example from a flotation concentrate) to provide an autogenous heat source, it is

impractical to achieve these pulp temperatures.

Beyond this, efficient methodologies, apart from conventional oxygenation, for oxidising the gold

tellurides at ambient temperature are less understood. Given that the gold tellurides oxidise

more rapidly at finer grind sizes, indicating that their oxidation is less time dependent but rather

reliant on oxygen mass transfer, then it is highly likely that if efficient methods of oxygen transfer

are applied, then the gold tellurides may be oxidised over shorter time intervals.

If this is the case and devices such as oxygen injectors, pipe reactors, etc, are applied, then

leach residues of 0.06 g/t Au would be achievable over time frames not dissimilar to

conventional plants, that is, 24 h oxidation and residence time.

Such residues equate to a percentage gold recovery of approximately 96% at a head grade of

1.47 g/t Au.

The key design criteria from a metallurgical perspective for each of the mineralogical domains

are shown in Table 13.7_1.



Table 13.7_1

Mara Rosa Project

Pre-Oxidation & Leach Design Criteria for Mara Rosa Samples

Criteria Unit

P80 grind size µm 45 pH set point 12.0 Pre-oxidation time h 12 Cyanidation leach time h 24 Cyanide consumption kg/t 0.26 Lime consumption kg/t 1.93 Gold head grade g/t Au 1.47 Gold in residue (design) g/t Au 0.10 Gold recovery (design) % 93.2 Gold in residue (optimum) g/t Au 0.06 Gold recovery (optimum) % 95.9

The details of gold recoveries for different grinds, pH, pre-oxidation times and leach times for

the individual leach tests are included in Appendix A of Smith and Witt (2011).

13.7.1 Grind Size

Pre-oxidation and cyanidation leach tests were initially conducted at a P80 grind size of 45 µm.

The relatively fine grind size was selected so as to afford the gold telluride species the best

opportunity for complete oxidation within the pre-oxidation and cyanidation leach time frames.

The tests carried out at a P80 grind size of 45 µm yielded good gold recoveries and so coarser

P80 grind sizes of 53 µm and 75 µm were also tested.

The testwork results displayed a strong correlation between grind size and gold recovery,

although this is less likely to be the result of gold liberation, as would normally be the case,

but instead a dependency on the ability of the gold telluride particles to oxidise during the pre-

oxidation and cyanidation leach stages.

The results at a P80 grind size of 53 µm and 75 µm were still reasonable although gold

recoveries decreased by between 2% and 4%.

Figure 13.7.1_1 shows the grind size – recovery relationship at pH 12 after 24 hours.



Figure 13.7.1_1

Grind Size Recovery Relationship at pH12 after 24 hours

88.0

89.0

90.0

91.0

92.0

93.0

94.0

95.0

96.0

40 45 50 55 60 65 70 75 80

Recovery (%)

Grind Size (um)

MAIN

HW

Based on the testwork results, a P80 grind size of 45 µm has been recommended, although it

should be noted that this grind size is based on laboratory screening, whilst an actual plant

would utilise hydrocyclones, which classify according to mass. The effect of this is that the

minerals with higher specific gravities, such as sulphides and tellurides, would be ground finer

due to their specific gravity and the consequent bimodal classification effect, meaning that a

P80 grind size of 53 µm would likely produce a P80 of 45 µm for the sulphide and telluride

particles.

13.7.2 Lime Demand and Effect of pH

The oxidation rate of gold tellurides is considerably slower at a pH of less than 12 and at

lower pulp temperatures (<60oC). Although it is impractical to raise the temperature, the pH

can be elevated by the addition of lime. The amount of lime required to raise the pH from

neutral (7.0) to 9.0 and then to 12.0 at increments of 0.5 was measured.

To achieve and maintain a target pH of 12.0 for 12 h pre-oxidation and 24 h of cyanidation

leach, 2.14 kg/t, 1.78 kg/t and 1.86 kg/t of lime were required for the main, hanging wall and

foot wall composite samples respectively.

The lime demand results for the main and hanging wall samples are shown in Figure 13.7.2_1

and Figure 13.7.2_2 respectively.



Figure 13.7.2_1

Main Composite Lime Demand

Figure 13.7.2_2

Hanging Wall Composite Lime Demand

Good gold recoveries were achieved at the initial target pH of 12 and so some additional

testwork was carried out at a pH of 11, however, the gold telluride oxidation rate decreased

significantly affecting overall gold recovery.



13.7.3 Pre-Oxidation and Cyanidation Time Leach

The results of the pre-oxidation and cyanidation time leach are discussed together as the exact

time required to oxidise the gold tellurides is difficult to ascertain and there is little doubt that the

gold telluride oxidation process is continuing to occur through the cyanidation leach tests.

Hence the separation of the pre-oxidation and cyanidation leach phases is somewhat

academic and for the purposes of telluride oxidation, should be combined, whilst for the

purposes of cyanidation leach, the important aspect is that non-telluride gold has sufficient

time to be dissolved in the cyanide solution.

A number of time combinations were tested for the main and hanging wall composite

samples. The aim was to test typical practical time frames as well as to provide ideal

maximum time frames, where complete gold telluride oxidation and gold cyanide dissolution

should have been possible.

13.7.4 Cyanidation Leach Residues

Detailed diagnostic analysis on the cyanidation leach residues was not required due to the

low gold residue values that were achieved.

13.8 Geochemistry

13.8.1 Introduction

Coffey were requested by Amarillo Gold to carry out geochemical testwork and analysis on

the leached tailings samples derived from the Mara Rosa metallurgical testwork.

The testwork focussed on:

Acid formation potential through ANC, NAG, NAPP testing (definitions below).

Multi-element composition of the tailings solids.

The geochemistry testwork was carried on two samples of CIL residue derived from previous

metallurgical testwork conducted at Ammtec Laboratories in Perth. The gold recovery

testwork was conducted on a number of composites samples representing the dominant

mineralogical lithologies present in the Posse Deposit.

Two leach residue samples from the ‘Main’ and ‘Hanging Wall’ composites were selected for

preliminary geochemical analysis as they form the majority of the mineralisation source. The

two samples represent the tail or residue from the pre-oxidation and cyanidation process.

The results indicate that both material types were ‘potentially acid forming’ (PAF), particularly

the ‘Main’ composite sample which was classified in the moderate to high range.



13.8.2 Samples

The residue samples used for geochemistry testwork were generated from gold leaching

testwork that was part of the preliminary metallurgical investigation conducted on the Posse

Deposit. It is assumed that the metallurgical testwork was conducted on representative feed

material and that the tails produced from the leaching testwork is indicative of the final tails

that could be expected from an operating plant.

13.8.3 Testwork Programme

The testwork procedures employed for this study are based on standard geochemical

characterisation methods. The testwork was completed by AMMTEC Ltd (AMMTEC) under

the project number A13025.

The two samples selected were the leach residues from cyanidation testwork samples:

HS24243 Main Composite

HS24244 Hanging Wall Composite

13.8.4 Acid-Base Chemistry

Maximum Potential Acidity (MPA)

The MPA reflects the maximum amount of acid that is generated if all the sulphide sulphur in

the sample is completely oxidised according to the following reaction:

FeS2 + 15/4O2 + 7/2H2O = Fe(OH)3 + 2 H2SO4

From the elemental analysis the sulphide sulphur grade of the residue samples were as

follows:

Main Composite: 1.46% 44.7 kg H2SO4 per tonne of mineralisation

Hanging Wall Composite: 0.82% 25.1 kg H2SO4 per tonne of mineralisation

Acid Neutralisation Capacity (ANC)

In this test the sample is acidified with a known amount of hydrochloric acid which is then

heated to ensure reaction completion. The calcium carbonate equivalent of the sample is

obtained by determining the amount of unconsumed acid by titration with standardised

sodium hydroxide.

The ANC results for the two tests were as follows:

Main Composite: 36.0 kg H2SO4 per tonne of mineralisation

Hanging Wall Composite: 44.6 kg H2SO4 per tonne of mineralisation



Net Acid Producing Potential (NAPP)

The NAPP is calculated from the corresponding MPA and ANC values as follows:

NAPP = MPA - ANC

The calculated NAPP for the Mara Rosa tails samples are indicated below. A negative value

for NAPP indicates that no acid should be produced. The acid neutralising component or

buffer potential is higher than the maximum amount of acid that could be produced:

Main Composite: = 44.7 – 36.0 = 8.7 kg H2SO4 per tonne of mineralisation

Hanging Wall Composite: = 25.1 – 44.6 = -19.5 kg H2SO4 per tonne of mineralisation

Net Acid Generating (NAG)

In this test the sample is placed under oxidising conditions to accelerate the sulphide

oxidation. The resulting solution is then back titrated to measure the amount of acid that was

produced.

The following results were obtained for the Mara Rosa Project samples:

Main Composite: 11.6 kg H2SO4 per tonne of mineralisation

Hanging Wall Composite: -13.3 kg H2SO4 per tonne of mineralisation

This is a physical test designed to validate the theoretical NAPP calculations above. It uses

accelerated oxidation to simulate the environmental conditions that occur in a tails storage or

waste dump environment over a long period of time.

Results Discussion

Table 13.8.4_1 presents a summary of the acid base testwork results.

Table 13.8.4_1

Mara Rosa Project

Acid Base Results Summary

Parameter Units Main Hanging Wall

Sulphide Sulphur % Sulphide Sulphur 1.46 0.82

MPA kg H2SO4 / tonne 44.7 25.1

ANC kg H2SO4 / tonne 36.0 44.6

NAPP kg H2SO4 / tonne 8.7 -19.5

NAG kg H2SO4 / tonne 11.6 -13.3

ANC/MPA ratio 0.81 1.78

The results above indicate the main wall composite has a relatively high potential for acid

formation. This is largely due to the high sulphide sulphur content of 1.46%. The acid

neutralising capacity of the material is also relatively high at 36.0 kg H2SO4 per tonne of



mineralisation however this is not enough to completely neutralise all the sulphuric acid that

could form, hence the positive NAPP reading.

The footwall samples had a lower sulphide sulphur content resulting in a significantly lower

MPA value of 25.1 kg H2SO4 per tonne of mineralisation. In this case however the material

had a very high buffering capacity with an ANC of 44.6 kg H2SO4 per tonne of mineralisation.

Hence the neutralisation capacity is higher than the potential acid formation which results in a

negative NAPP value.

The net acid generation or NAG test resulted in a good correlation with the theoretical data

with the Main sample producing 11.6 kg H2SO4 per tonne of mineralisation and the Hanging

Wall producing -13.3 kg H2SO4 per tonne of mineralisation.

In recent years, research (especially estimation of reaction-rates for diverse sulphide/gangue

mineral assemblages) and field-experience at mining operations world-wide have shown that

the potential for acid mine drainage (AMD) production is very low for mineralised waste and

tailings materials with ANC/MPA ratios greater than 2.0.

The acid forming potential (AFP) of a sample can be classified as either:

Non-Acid forming (NAF)

Potentially acid forming (PAF)

The classification criteria often used in mining operations worldwide are:

NAF: Sulphide Sulphur <0.3%, both a negative NAPP and an ANC/MPA ratio of ≥2.0

PAF: Sulphide Sulphur ≥0.3%, a positive NAPP or a negative NAPP value with an

ANC/MPA ratio of <2.0

Based on the results obtained the Main composite would be classified as PAF and is in the

moderate to high range for tailings materials. The high NAPP and the confirmatory NAG test

indicate that this material is likely to generate a significant amount of acid in an oxidising

environment such as tail facilities or waste dump.

The Hanging Wall sample appears to be more ‘benign’ mainly due to the lower sulphide

sulphur content. Overall the material produced a negative NAPP value which was confirmed

with the negative NAG result.

Whilst most indications point to the hanging wall material not being an acid producer,

according to established guidelines, material with ANC/MPA ratios less than 2.0 are

technically classified as ‘potentially acid forming’. In this case the ANC/MPA ratio is slightly

low at 1.78, however based on the overall results it would appear that the risk of significant

acid formation associated with this material is low.



13.8.5 Multi Element Analysis

Tails Solids

The multi-elemental analysis of the tailing samples are presented in Table 13.8.5.1_1 and

13.8.5.1_2 for the Main and Hanging Wall samples respectively, along with a comparison with

the average crustal abundance of the earth and the Geochemical Abundance Index (GAI). The

GAI is calculated from the ratio of the sample element content and the average crustal

abundance.

A GAI greater than 3 usually signifies enrichment to a level that warrants further investigation.

Element enrichments serve as a starting point in the assessment of potential concerns for

element leaching and the production of toxic dust from dry exposed tailings in the storage

facility.

Both samples are similar in elemental composition with the exception of antimony which is

significantly higher in the hanging wall sample. Results indicate that both molybdenum and

titanium may be enriched in both tails materials. Antimony is considered slightly enriched in

the hanging wall sample.

Table 13.8.5.1_1

Mara Rosa Project

Tails Multi Elemental Analysis – Main Composite

Element Units Element Content Average Crustal

Abundance (ACA) Geochemical

Abundance Index (GAI)

Al % 7.76% 8.20% 0 Ca % 2.10% 4.10% 0 Fe % 3.70% 4.10% 0 K % 2.33% 2.10% 0 Mg % 0.84% 2.30% 0 Na % 3.26% 2.30% 0 As ppm 10 1.5 2 Ba ppm 600 500 0 Cd ppm 0 0.11 0 Co ppm 40 20 0 Cr ppm 1100 100 2 Cu ppm 240 50 1 Hg ppm 0.1 0.05 0 Mn ppm 825 950 0 Mo ppm 120 1.5 5 Ni ppm 605 80 2 P ppm 700 1000 0 Pb ppm 10 14 0 Sb ppm 0.3 0.2 0 Se ppm 0 0.05 0 Sr ppm 156 370 0 Th ppm 14 12 0 Ti ppm 3400 0.6 6 U ppm 4.4 2.4 0 V ppm 74 160 0 Zn ppm 96 75 0



Table 13.8.5.1_2

Mara Rosa Project

Tails Multi Elemental Analysis – Hanging Wall Composite

Element Units Element ContentAverage Crustal

Abundance (ACA) Geochemical

Abundance Index (GAI)

Al % 7.84% 8.20% 0 Ca % 2.61% 4.10% 0 Fe % 3.98% 4.10% 0 K % 3.03% 2.10% 0 Mg % 1.44% 2.30% 0 Na % 3.46% 2.30% 0 As ppm 10.00 1.5 2 Ba ppm 1200 500 0 Cd ppm 0 0.11 0 Co ppm 40 20 0 Cr ppm 1100 100 2 Cu ppm 180 50 1 Hg ppm 0.2 0.05 1 Mn ppm 985 950 0 Mo ppm 90 1.5 5 Ni ppm 565 80 2 P ppm 700 1000 0 Pb ppm 15 14 0 Sb ppm 5.9 0.2 4 Se ppm 0 0.05 0 Sr ppm 246 370 0 Th ppm 12 12 0 Ti ppm 3400 0.6 6 U ppm 4 2.4 0 V ppm 88 160 0 Zn ppm 86 75 0

Enriched minerals are only considered to be problematic if they are soluble and may

potentially be leached in the tails environment. Leaching and mobilisation of the metals may

occur under acid conditions that can results from acid mine drainage environments.

Molybdenum and titanium are not considered problematic and will typically form stable oxides in

a tails environment. Antimony is relatively stable and forms various salts and is not typically

reactive in dilute acidic or alkali environments. Antimony is toxic however and could potentially

become problematic in the form of dust generated from a dry tails facility for example.

13.8.6 Conclusions

Based on the preliminary testwork results obtained in this study it would appear that the risk of

significant acid formation associated with the main composite material from the Posse Deposit

is relatively high. The hanging wall material which forms the other major component of the

mineralisation however was relatively benign.

Both material types would be classified as PAF according to the industry standard guidelines

however the hanging wall material is considered a low risk in terms of acid production

potential.

Further investigation is required in order to determine a suitable tails management strategy in

order to mitigate the potential acid formation in a tails facility. The management plan could



range from simply blending the high and low acid sources to minimise the risk, to engineering

solutions incorporated into the process plant and tails facility designs.

It is also recommended that samples of the Mara Rosa waste rock material be sent for

geochemical testing to assess the potential for acid formation in waste dumps.

The elemental results indicated that a titanium and molybdenum metals were enriched in both

of the residue samples however these are relatively stable and not considered problematic in

a tails environment. The hanging wall material had significantly higher levels of antimony than

the main composite sample which can be toxic and hence problematic, particularly in the form

of dust that can be generated from dried tailings. The antimony content is however based on

one analysis and further testwork would be required to confirm the result.

13.9 Ancillary Testwork

No detailed ancillary testwork was undertaken at this stage, although the following comments

are relevant to the progression to more detailed levels of study.

13.9.1 Site Water

The testwork was completed using Perth tap water. The effect of this is considered negligible

as the site water is of a high quality and contains no deleterious elements or minerals that

would affect the gold recovery for the proposed process flowsheet. It should be noted that

site water analysis has been completed as a part of the environmental baseline studies.

13.9.2 Settling / Thickening

Settling / thickening testwork was not completed at this point and would not be expected to be

an issue. It is proposed that a thickener precede the cyanide detoxification step, so as to

recover cyanide and minimise the detoxification requirement. Given the fresh non viscous

nature of the deposit, a conservative settling rate of 0.50 t/m²/h has been applied for the

thickener design. Despite this, such testwork should be completed as part of the feasibility

stage of testwork.

13.9.3 Pulp Viscosity

The samples displayed no viscosity issues during the leach testwork and hence agitator

design is not considered an issue even at a P80 grind of 45 µm. Despite this, such testwork

should be completed as part of the feasibility stage of testwork.



13.10 Key Design Criteria

The summary key design criteria is shown in Table 13.10_1.

Table 13.10_1

Mara Rosa Project

Design Criteria for Mara Rosa Samples

Criteria Unit

Annual throughput tpa 2,500,000

Availability % 90.0%

Instantaneous throughput tph 317

Bond rod mill work index kWh/t 13.4

Bond ball mill work index kWh/t 13.0

Abrasion index 0.3426

P80 grind size µm 45

pH set point 12.0

Pre-oxidation time h 12

Cyanidation leach time h 24

Cyanide consumption kg/t 0.26

Lime consumption kg/t 1.93

Thickener settling rate t/m²/h 0.50

Gold head grade g/t Au 1.47

Gold in residue (design) g/t Au 0.10

Gold recovery (design) % 93.2

Gold in residue (optimum) g/t Au 0.06

Gold recovery (optimum) % 95.9

13.11 Process Flowsheet

The Mara Rosa metallurgical samples were comprised of a free milling component of gold

and a refractory component which may be associated with sulphides as well as tellurides.

The gold can be recovered using conventional CIL techniques, albeit with a requirement for

finer grinding and pre-oxidation prior to leaching.

It is proposed that the mineralisation would undergo primary crushing followed by secondary

and tertiary crushing in closed circuit. Tertiary crushed material would then feed a primary mill

and secondary mill utilising cyclone classification to achieve a P80 grind of approximately 45 µm.

This material would be pre-oxidised at a pH of 12 for a period of 12 hours in agitated tanks

and then leached under conventional CIL conditions for 24 hours during which an average of

93% of the gold would be dissolved and adsorbed onto activated carbon. The adsorbed gold

would then be eluted using a conventional desorption plant.

Tailings would be thickened to recover a significant proportion of the process solution (pH 12

and CN 0.015%), with the remaining thickened pulp detoxified to remove residual free cyanide

prior to deposition in a tailings storage facility.



There is considerable scope to improve the pre-oxidation stage of the process, which

currently accounts for 12 hours of agitated tank residence time. This may be achieved by

oxygen injectors, pipe reactors, or other innovative means.

13.12 Further Testwork Recommendations

Further testwork is deemed necessary assuming the project proceeds to a feasibility level of

study. The testwork required at this stage would need to include variability testwork, as well

as confirmatory composite testwork and optimisation testwork on the grind versus pre-

oxidation characteristics of the mineralisation.

In addition to this, ancillary testwork requirements as discussed in Section 13.9, would need

to be completed.

The details of the required testwork programme are included in Appendix A of Smith and Witt

(2011).



14 MINERAL RESOURCE ESTIMATES

AEFS completed a Mineral Resource Estimate in July 2011 using drillhole details supplied by

Amarillo and wireframes supplied by HCS. The Mineral Resource Estimation presented below

is based upon historic drill holes included as part of the previous resource estimate, together

with the results from a series of 33 diamond drill holes drilled between late 2010 and mid 2011

discussed in Section 10.

This estimate represents the fourth Independent Mineral Resource Estimate completed on

behalf of Amarillo for the Posse Deposit. CCIC completed estimation of an Inferred Mineral

Resource Estimate in March, 2007. An updated resource estimate which complied with the

“Canadian Institute of Mining, Metallurgy and Petroleum Standards on Mineral Resources and

Mineral Reserves Definitions Guidelines” was provided to Amarillo by CCIC in February 2008

while HCS and AEFS completed a Mineral Resource Estimate in 2010.

14.1 Data Utilised

Data used in the modelling consisted of a Micromine format set of drill collars, downhole

surveys and assay data together with a set of wireframe boundaries which defined Hanging

Wall, Main and Foot Wall zones to the deposit and the current surface topography. The

wireframes were developed by Hugo Hoogvliet of HCS using the latest set of drillhole data.

The surface topography used reflects the work discussed in section 12.4 and reflects the

historic pits mined by WMC. A raw block model with primary blocks of 25 (E), 25 (N) and 10

(RL) meters which was sub-blocked to 5 m x 5 m x 5 m to fit the ore and topography

wireframes was also generated. The blocks in this model were coded to indicate which

portion of the ore body a block belonged to. All data had coordinates in terms of the WGS 84

coordinate system.

A default SG of 2.73 (the same value as used in the 2010 report) was used for all tonnage

calculations. Further SG measurements were made at site using the drill core from the most

recent drilling which suggest that the density of 2.73 may need to be revised upwards,

however this data had not been validated prior to the development of this resource estimate

so was not used.

The modelling ellipses (with minor adjustments to improve the orientation of the ellipses)

together with the variography used in the 2010 model were used in the 2011 model.

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

14.2 The Modelling Process

Initial modelling work consisted of defining a set of consistent sections to be used in the

development of wireframes. The sections were all at an angle to the grid so that they are at

right angles to the strike of the mineralisation.



Three grade zones were recognised, Hanging Wall, Main and Foot Wall. The grade in both

the Hanging Wall and Foot Wall was lower than the grade in the Main Zone and neither the

Hanging Wall or Foot Wall zones are continuous. There are also occasional spurs off the

Hanging Wall, these have not been modelled as they are not well defined by drilling, similarly

the basaltic dyke discussed in Section 7.3 above has not been incorporated in the wireframe

as there is currently insufficient information to define its spatial orientation. During

construction of the wireframes the boundary of the Hanging Wall and Foot Wall was

deliberately allowed to intrude into the Main zone. After the initial wireframes were built they

were then clipped using the Wireframe Boolean functions in Micromine to produce wireframe

that had no under or overlap. The wireframes are shown in Figure 14.2_1.

Figure 14.2_1

Posse Wireframes and drillholes

Assay data was coded according to which wireframe the assay related to and was then

composited to 1 m intervals for modelling. This had the effect of reducing high grades and no

further cutting of the grade was used. Key statistical parameters of the raw and composited

data are shown in Tables 14.2_1 and 14.2_2.



Table 14.2_1

Mara Rosa Project

Summary statistics for raw assays in the mineralised zone

Raw Assays

All Zones HW Main FW

NORMAL STATISTICS

Minimum 0.00 0.00 0.00 0.01

Maximum 90.00 19.00 90.00 12.20

No of points 23168 3165 5552 1001

Mean 0.57 0.51 1.73 0.50

Variance 3.07 1.06 9.26 0.50

Standard deviation 1.75 1.03 3.04 0.71

Median 0.13 0.30 0.95 0.35

Coefficient of variation 3.10 2.02 1.76 1.43


LOGARITHMIC STATISTICS

No of points 23168 3165 5552 1001

Mean of natural logs -2.13 -1.18 -0.00 -1.12

Geometric Mean 0.12 0.31 1.00 0.33

Geometric Std dev 6.42 2.46 2.70 2.44

Natural Log variance 3.46 0.81 0.99 0.79

Nat Log Std veviation 1.86 0.90 0.99 0.89

Sichel's V 3.46 0.81 0.99 0.80

Sichel's Gamma 5.63 1.50 1.64 1.49

Sichel's T-Estimator 0.67 0.46 1.64 0.49



Table 14.2_2

Mara Rosa Project

Summary statistics for 1m composite assays in the mineralised zone

1m Composites


NORMAL STATISTICS

Minimum 0.01 0.00 0.00 0.00

Maximum 41.17 19.00 41.17 10.65

No of points 9396 3197 5219 1035

Mean 1.17 0.51 1.69 0.49

Variance 3.97 0.98 5.85 0.40

Standard deviation 1.99 0.99 2.42 0.63

Median 0.60 0.30 1.01 0.33

Coefficient of variation 1.71 1.95 1.43 1.30


LOGARITHMIC STATISTICS

No of points 9396 3168 5201 1027

Mean of natural logs -0.47 -1.14 0.06 -1.09

Geometric Mean 0.62 0.32 1.06 0.34

Geometric Std dev 2.93 2.35 2.53 2.34

Natural Log variance 1.15 0.73 0.86 0.72

Nat Log Std veviation 1.07 0.85 0.93 0.85

Sichel's V 1.15 0.73 0.86 0.72

Sichel's Gamma 1.78 1.44 1.54 1.44

Sichel's T-Estimator 1.11 0.46 1.64 0.48

The block model was based on 25 m x 25 m x 10 m blocks sub blocked to 5 m x 5 m x 2 m to

ensure that blocks fitted the wireframe boundaries. Each block was coded to identify the

wireframe each block belonged to and to identify if the block was above or below the

topography. The model included blocks above the current topography as the dataset being

modelled included a number of holes drilled from inside the former open pit. The wireframes

were constructed to include this data however blocks above the current surface were

removed from the model.

The modelling process consisted of four (4) runs through the data for each of the zones to be

modelled (HW, Main and FW) using different sized search ellipsoids. The different ellipsoids

were nominally used to differentiate which resource classification code would be assigned to

a block (i.e. Measured, Indicated, Inferred and Not Classified) this process was then modified

based on the drillhole data density as shown on a long section in the plane of the vein, (Figure

14.3_1).

The data was modelled using Median Indicator Kriging and had a grade that was in keeping

with the expected grade and showed a good fit to the raw data. The modelling result, while



inevitably over estimating lower grades and underestimating higher grade, a result of the

volume variance effect, should still reflect the distribution of the raw data.

The search ellipsoid used in the modelling process was based on that used in the 2010

modelling exercise. The orientation of the search ellipsoid was however updated to provide a

better fit with the model anisotropy defined by the variography. Note that as the median value,

the indicator modelled remained the same as in the 2010 model and there was no need to

update the 2010 variography shown in Figures 14.2_2.

Figure 14.2_2

Downhole and Directional semi-variogram, Au median indicator

Donwhole

Along strike

Down dip Across dip

The search parameters outlined in Tables 14.2_3 and 14.2_4 were used in the model for each

of the domains, HW, Main and FW :



Table 14.2_3

Mara Rosa Project

Key modelling parameters

Run Long Axis Intermediate Axis Short Axis Points References

Length Azimuth Plunge Length Dip Length Minimum Minimum

1 35 247 21 14 38.21 3.5 4 4

2 70 247 21 28 38.21 7 2 2

3 140 247 21 56 38.21 14 2 2

4 280 247 21 112 38.21 28 1 2

Table 14.2_4

Mara Rosa Project

Modelling parameters common to all modelling runs

Sectors References

Number Max points Minimum Count Maximum Count

8 10 1 10

The median indicator kriging used bins based on percentiles of the data as shown in Table

14.2_5.

Table 14.2_5

Mara Rosa Project

Modelling parameters common to all modelling runs

Bin Cutoff Percentile

1 0.03 1.0

2 0.10 5.0

3 0.15 10.0

4 0.25 20.0

5 0.30 25.0

6 0.33 30.0

7 0.45 40.0

8 0.60 50.0

9 0.77 60.0

10 1.04 70.0

11 1.20 75.0

12 1.49 80.0

13 1.85 85.0

14 2.50 90.0

15 4.17 95.0

16 6.2 97,.5

17 10.2 99.0



The final MIK model provided a reasonable estimate of the grade and has also minimised

overestimation of low grades zones and underestimation of high grade as shown in Figure

14.2_3. Key statistics comparing the input data and the modelled data are setout in Table

14.2_6.

Figure 14.2_3

Comparison between 1m Composites and grades in the merged MIK model

Table 14.2_6

Mara Rosa Project

Key model statistics compared to the input data

Statistic 1m composite; All zones Model; All zones

Mean 1.16 1.11

SD 1.99 1.12

Median 0.6 0.71

CC (normal) 0.95

CC (ln) 0.99

14.3 Model Classification

After generating the block model the blocks within the model were classified. This was done

by visually assessing the density of drillholes that pierced the ore zone using an inclined long

section. The classifications applied to different zones within the mineralised horizon are

shown in Figures 14.3_1 below.



Figure 14.3_1

Comparison between 1m Composites and grades in the merged MIK model

The areas of the model enclosed by the red boundary were classified as Measured, Those

enclosed by the Blue boundary as Indicated and those enclosed by the green boundary as

Inferred. Any blocks falling outside the green boundary have been classified as NC, Not

Classified, and have not been reported.

14.4 Resource

The resource above a cutoff of 0.5 g/t declared for the Posse Deposit is summarised in the

Table 14.4_1 while a grade tonnage curve for the deposit is shown in Figure 14.4_1.



Table 14.4_1

Mara Rosa Project

Resource Summary

Grade and Tonnes Above Specified Cutoff

From Volume Tonnes Au (g/t) Metal (oz) % Total Domain Material

0.5 357,000 973,000 0.68 21,300 14 HW Inferred

0.5 1,602,000 4,373,000 1.1 154,700 19 HW Indicated

0.5 319,000 871,000 0.71 19,900 6 HW Measured

0.5 2,277,000 6,217,000 0.98 195,900 15 HW Total

0.5 777,000 2,121,000 1.81 123,400 79 MAIN Inferred

0.5 3,685,000 10,060,000 1.96 633,900 78 MAIN Indicated

0.5 1,537,000 4,195,000 2.44 329,100 92 MAIN Measured

0.5 5,999,000 16,377,000 2.07 1,089,900 82 MAIN Total

0.5 196,000 535,000 0.67 11,500 7 FW Inferred

0.5 352,000 960,000 0.81 25,000 3 FW Indicated

0.5 145,000 396,000 0.78 9,900 3 FW Measured

0.5 693,000 1,892,000 0.76 46,200 3 FW Total

0.5 1,330,000 3,630,000 1.34 156,400 Total Inferred

0.5 5,638,000 15,393,000 1.65 816,600 Total Indicated

0.5 2,001,000 5,463,000 2.04 358,300 Total Measured

0.5 8,969,000 24,485,000 1.69 1,330,400 Total Total

Note: Mineral Resources are estimated by Keith Whitehouse, MAusIMM (CP), QP, of

Australia and have an effective date of 31 May, 2011. Mineral Resources that are not Mineral

Reserves do not have demonstrated economic viability. Mineral Resources are inclusive of

Mineral Reserves. Tonnages are metric tonnes and ounces of contained gold are troy ounces.

Mineral Resources above a 0.5 g/t Au cutoff grade have reasonable prospects for economic

extraction, based on mineralization continuity, shape and distribution and as demonstrated in

this study.



Figure 14.4_1 Grade Tonnage Curve



15 MINERAL RESERVE ESTIMATES

Table 15_1 shows the Mineral Reserve estimate, based on a Mineral Resource cutoff grade

of 0.5 g/t Au. The Mineral Reserve is included within the declared Measured and Indicated

Mineral Resource and is declared inclusive of approximately 0.5 Mt of dilution at an average

grade of <0.2 g/t.

Table 15_1

Mara Rosa Project

Mineral Reserve Estimate (28 October 2011)

Classification Tonnes (t) Au grade (g/t) Contained Gold (oz)

Proven Mineral Reserve 5,366,400 1.97 339,600

Probable Mineral Reserve 11,750,400 1.60 606,600

Total Mineral Reserve 17,116,800 1.72 945,200 The tonnes and grade reported here is Run of Mine. Application of the plant recovery factor reduces the recoverable

gold to 869,600 oz.

Rounding has been applied.

The Mineral Reserve estimate has been determined and reported in accordance with the CIM

Definition Standards (2010).

The reported Mineral Reserve has been compiled under the supervision of João Augusto

Hilário, BSc, MAIG, an employee of Coffey Consultoria e Serviços Ltda.

A summary of the main input factors used in estimating the Mineral Reserve are shown in Table 15_2.

Table 15_2

Mara Rosa Project

Input Parameters used for the Mineral Reserve Estimate (28 October 2011)

Description Units Value

Gold price US$/oz 1,100 Mineral resource Au cut off grade g/t 0.5 Mining method Open pit Annual production rate Mtpa 2.5 Mining operating cost US$/t ore 12.59 * Processing operating cost US$/t ore 9.73 ** G&A operating cost US$/t ore 1.83 Mining dilution % 3 Mining recovery/loss % 97 Plant recovery % 92 Project capital cost US$M 184 Sustaining capital cost US$M 11 Royalty % 2 Pit slope degrees 55° HW 40° FW Strip ratio 8:1 * Mining operating costs are quoted in this table inclusive of the Year 0 pre-stripping.

** Processing operating costs are estimated for a plant design throughput of 2.5 Mtpa.



The data that supports the Mineral Reserve is discussed in other sections of this technical

report and was obtained from sources listed in Table 15_3.

The Mineral Reserve estimates were constrained by mining, metallurgical and infrastructure

factors as summarised here, and detailed elsewhere in this report. The reserves are not

expected to be affected by permitting or any other factors.

Table 15_3

Mara Rosa Project

Mineral Reserve Estimate Sources of Supporting Information (28 October 2011)

Modifying Factor Source

Mineral Resources AEFS/HCS

Geotechnical Engineering Coffey Mining

Mine Design Coffey Mining

Mine Cost Estimation Coffey Mining

Hydrology and Hydrogeology Hidrovia

Metallurgical Testwork Coffey Mining

Process Design Coffey Mining/Amarillo/Onix

Process Plant Design & Cost Estimate Amarillo/Onix

Infrastructure Design & Cost Estimate Coffey Mining/Onix

Tailings Storage Facility Coffey Mining

Environmental Neotropica

Social Neotropica

Marketing Coffey Mining

Financial Modelling Coffey Mining

Property and Land Tenure

A Mineral Resource cutoff grade of 0.5 g/t Au was used for this Pre-Feasibility Study and this

was based on the work completed for the Independent Mineral Resource Estimate and

Preliminary Economic Assessment (2010), the Report on Independent Site Visit and

Resource Estimate (2011), and a Whittle pit base Project gold price of US$1,100/oz. Table

15_4 shows the Project in situ breakeven Au grade based on operating costs and recoveries

estimated during the completion of this Pre-Feasibility Study for a range of gold prices up to

the 2011 peak price of US$1,900/oz.



Table 15_4

Mara Rosa Project

Break Even Grade Estimate

Parameter Gold Price (US$/oz)

Gold Price (US$/oz) 1,000 1,100* 1,200 1,400 1,600 1,800 1,900

Plant Recovery (%) 92 92 92 92 92 92 92

Mining Recovery (%) 97 97 97 97 97 97 97

Mining Cost (US$/t ore) 12.59 12.59 12.59 12.59 12.59 12.59 12.59

Plant Cost (US$/t ore) 9.73 9.73 9.73 9.73 9.73 9.73 9.73 G&A Cost (US$/t ore) 1.83 1.83 1.83 1.83 1.83 1.83 1.83

CFEM (US$/t ore) 0.56 0.56 0.56 0.56 0.56 0.56 0.56

Royalties (US$/t ore) 1.12 1.12 1.12 1.12 1.12 1.12 1.12

Transport & selling cost (US$/t ore) 1.17 1.17 1.17 1.17 1.17 1.17 1.17

Total Cost (US$/t) 27.00 27.00 27.00 27.00 27.00 27.00 27.00

In Situ break even grade Au g/t 0.84 0.76 0.70 0.60 0.52 0.47 0.44

*Base case gold price



16 MINING METHODS

The following activities were undertaken by Coffey Mining:

Study of geotechnical design parameters provided by the BVP Engineering and reviewed

by Coffey Mining;

Open pit optimisation based on the resource block model provided by AEFS, and using

techno-economic parameters agreed between Amarillo and Coffey Mining;

Open pit design utilising medium size equipment appropriate to the scale of operation,

mine life and safe operation;

Development of the mining sequence and scheduling;

Planning of waste rock dumps, ore and low grade stockpiles;

Selection of tailings storage facility location, design and closure concept;

Development of the mine production and operating parameters;

Selection of mining equipment and estimation of productivity;

Description of pit operations and infrastructure.

16.1 Geotechnical

16.1.1 Previous studies

Caracle Creek International Consulting

A report on the Mara Rosa property was compiled by Caracle Creek International Consulting

Inc. in February 2008, from which the general description and geological summary above are

taken. The report describes the general site conditions, tenement conditions and ownership

history, regional and deposit geology, mining history, investigative works undertaken to date

and economic and environmental considerations.

Hoogvliet Contract Services & Australian Exploration Field Services Pty Ltd

A further mineral resource estimate and preliminary economic assessment was carried out by

Hoogvliet Contract Services and Australian Exploration Field Services Pty Ltd in June 2010.

The report summarises general site conditions, work done to date, tenement history, regional

and deposit geology, metallurgical test work and other relevant considerations, in addition to

providing an updated mineral resource estimate.

BVP Assessment

The assessment carried out by BVP in February 2011 was based on a site visit, during which

two BVP engineers examined the rock mass exposed in Cava Norte and Cava Sul pits,

available drill core and samples, and a range of documents including the two reports



described above. The main technical aspects of BVP’s assessment are summarised in the

following sub-sections.

Geotechnical Logs of Drill Core

The geotechnical logs of the bore holes from the two phases of ground investigation are

broadly similar, although there are differences, particularly in the way RQD is assessed. In

one of the phases, it was assessed from 12.5 cm lengths of intact core, rather than the normal

10cm lengths stipulated by Deere and Deere (1988)1, who proposed the RQD concept. The

rock strengths recorded from the drill core were assessed to be somewhat conservative by

BVP; following their review of the logs and assessment of the core which was examined on

site. BVP recommended that the rock strengths logged might be more appropriately

assessed as being in the strong to very strong range (R4-R5), rather than the moderately

strong to strong (R3-R4) range interpreted by Amarillo loggers. In all other respects, the

logging was assessed by BVP to be of an appropriate standard. BVP assessed the rock mass

in general to be of good to very good quality, although this assessment is empirical rather

than being based on an assessment of Rock Mass Rating (RMR).



Structural Interpretation

As the core was not oriented, no data on orientations of specific discontinuities within the core

is available. Field observations and examination of geological maps led BVP to conclude that

schistose/gneissose fabric is the dominant pervasive structural feature, dipping at between

40° and 55° to the north-west. The presence of a major structure located on the current

hangingwall of Cava Sul Pit with a long persistence, dipping steeply (estimated to be greater

than 65°) toward the south-east, was also noted by BVP during the site visit.

Preliminary Pit Slope Design

BVP suggest that, for preliminary slope design purposes, the footwall slope angle should

follow the dip of the orebody, which is generally parallel to the foliation, resulting in an angle of

40° to 45°. In the weathered part of the footwall, it is stated that a lower angle should be

adopted. An overall angle in the order of 55° is suggested for the hangingwall, based on

experience of other mine slopes in similar rocks. It is emphasised that these angles are

preliminary, and are subject to suitable blasting techniques being adopted which do not

compromise the integrity of the rock mass.

Review of BVP report

At the request of Amarillo, Coffey Mining undertook a review of BVP’s report in March 2011.

Discussions were held with BVP’s engineers on one or two outstanding queries; as a result of

which it was Coffey Mining’s assessment that BVP had appropriately summarised the

information made available to it during their site visit. Coffey Mining is in agreement with BVP

that sufficient information was provided for slope angles to be assessed for preliminary design

purposes.

As part of this report, Coffey Mining has independently assessed the overall footwall slope

angle for preliminary design purposes, using the Haines and Terbrugge empirical method,

which is described in Section 16.1.5. Coffey Mining concurs with BVP’s assessment that the

footwall should generally follow the dip of the pervasive foliated rock fabric, which

approximately parallels the mineralisation boundaries. However, assessment of the foliation

plane shear strength will need to be undertaken to determine whether the overall slope can be

cut parallel to the foliation angle, or whether the batter slopes would have to be cut parallel to

this (to avoid undercutting foliation at the batter-scale).

16.1.2 Deposit Geology

Solid Geology

Two main lithological units can be distinguished in the area of the Posse North and

Posse South Deposits. These are: fine grained gneiss, which forms the hangingwall of the

deposit, and; quartz-mica schist with intercalations of amphibolite, which forms the

mineralisation host and footwall. These units are within a ductile shear zone which strikes

northeast-southwest and dips between 45° and 55° to the north-west. Sections through the



Posse South Deposit (provided by Amarillo), together with a plot of drillhole and section line

locations are shown in the Beer (2011).

Weathering

The depth of weathering varies between the hangingwall and footwall. On the hangingwall,

weathering was noted by BVP to extend down to between 5 m and 15 m. In the footwall, the

weathering was noted to extend to a minimum of around 30 m depth.

Structures

The major structure within the deposit is the ductile shear zone which strikes northeast-

southwest and dips between 40° and 55° to the north-west.

A sub-vertical structure with a north-easterly strike was also noted on the south-east wall of

Cava Sul.

Finally, a large scale structure which dips at 65° or greater (the dip angle was estimated by

Coffey Mining from a photograph in the BVP report), was noted in the hangingwall outcrop by

BVP during their site visit. Although this does not appear to have a deleterious effect on the

stability of the hangingwall of the existing flooded pit, it is not known whether additional

structures of a similar orientation with lower shear strengths may also exist within the

hangingwall.

Minor structures (jointing) within the rock mass have not been assessed, as no oriented core

logging has been undertaken to date; however the hangingwall rock mass was characterised

by BVP as being ‘very little fractured’.

16.1.3 Geotechnical Data

Drillhole data

Two phases of geotechnical drilling have been carried out, as previously mentioned. These

can be distinguished by their prefix letters. The drillholes from the earlier phase are identified

by the ‘SPETI’ prefix and the later (and current) phase is identified by the ‘MRP’ prefix. The

location of the bore holes, Field Index Strength (FSI) and RQD values are shown in Figures

A8 to A10. No FSI values are shown for the SPETI holes as they are ranked ‘S’ or ‘M’; the

correlation with standard FSI is unknown. It should also be noted that the RQD values for the

MRP holes are based on the percentage length of core present in a minimum of 12.5cm long

intact core pieces; the standard length used to calculate RQD is 10 cm. The MRP RQD

values have been plotted as supplied as, at worst, they will represent conservative values for

RQD; the ‘standard’ values would be higher.

The histogram for the Field Index Strength values is shown below (total metres logged for

each strength class is shown) in Figure 16.1.3_1. The length weighted ‘average’ value for

Field Index Strength is R5.



Figure 16.1.3_1 Distribution of Field Strength Index values for Mara Rosa Project

The histogram for the RQD values is shown below in Figure 16.1.3_2. The length-weighted

average value for RQD is 79.7%; however if all values in the weathered zone are ignored, a

length weighted average of 82% is obtained.



Figure 16.1.3_2 Distribution of RQD Values for Mara Rosa Project

.

Field observations

Field observations made by BVP during their site visit suggested that the depth of weathering

on the hangingwall of the Cava Sul pit was very shallow (less than 5 m). The weathered

profile on the footwall (amphibolite) was deeper, extending to below the current water level in

the pit.

The field observations made of the hangingwall of the Cava Sul Pit confirmed the drillhole

data which suggests that the hangingwall is relatively massive, with little fracturing. It also

noted the presence of a persistent structure, striking parallel to the pit wall and dipping at an

angle estimated (by Coffey Mining from the BVP report photos) to be approximately 65° to the

south-east.

16.1.4 Geotechnical Model

Lithology and Weathering Distribution

Two major lithological units have been identified in the Posse Deposit:

Muscovite-biotite schist, with intercalations of amphibolite, which forms the footwall to the

deposit.

Fine-grained feldspar-biotite gneiss which forms the hangingwall to the deposit.



No detailed description has been provided to Coffey Mining of these lithologies, but the field

strength index, alteration and weathering have been described in the logs provided in the

geotechnical database.

These units are interpreted to dip at between 40° and 50° towards approximately 320° (UTM

Grid). Interpreted geological sections supplied by Amarillo are shown in Beer (2011). Other

minor lithologies have been recorded (e.g., a silicified zone, and a mafic tuff). These are

interpreted to be conformable with the major lithological boundaries.

From the drillhole data and field observations it has been assessed that the weathering depth

on the hangingwall gneisses is around 5 m, while the depth of weathering in the footwall

amphibolites may extend to around 30 m below surface.

Rock Material Characterisation

Field Strength Index

The Field Strength Index ratings for the drillhole database range from R0 to R6. Careful

examination of the database shows that the R0 ratings come solely from the soil horizon, and

can be discounted when forming strength models for the hangingwall and footwall. The most

common recorded ratings are R4-R6, and a length weighted ‘averaging’ approach suggests

that an R5 FSI may be appropriately adopted to characterise the hangingwall and mineralised

zone. (This implies a UCS strength of greater than 100 MPa for the intact fresh rock).

The footwall amphibolites are not extensively penetrated by the exploration drillholes, but

examination of the drill core that does comprise amphibolite suggests that it may be

appropriate to adopt an R4 or R5 rating for the footwall amphibolites. However, the

amphibolite is highly foliated, and it is anticipated that the shear strength along foliation will be

a more critical controlling factor on footwall stability than the compressive strength of the

intact rock.

Alteration and weathering

Little alteration or weathering were noted for the hangingwall either in the drill core or from

review of the field exposures by BVP. It is assessed that it would be appropriate to consider

the hangingwall to be within fresh rock from 5m below surface for the purposes of the PFS.

The footwall amphibolites are more extensively weathered than the hangingwall gneiss, with

the depth of weathering perhaps extending as far as 30 m below surface. As the first effects

of weathering are generally noted along discontinuities, it is assessed that, depending on the

degree of weathering, the discontinuity shear strength may be significantly reduced in the

footwall weathered zone.



Rock Mass Characterisation

RQD

A length-weighted average RQD value of 82% was obtained for the drill core from below the

weathered zone of the hangingwall. Examination of core on-site and comparison with the field

exposure suggests that it may be appropriate to adopt this as a conservative value for

assessment of Mining Rock Mass Rating for use in empirical slope design calculations. The

field exposure suggests that the fresh rock mass has very little fracturing, and that an RQD of

90% or higher may be appropriate; it is likely that many of the discontinuities taken into

account during core logging may in fact have been drill-induced fractures. It was noted during

core examination that the drillers’ marking of drilling-induced fractures was absent, which

would exacerbate this.

RQD has not been assessed for the footwall due to the very limited length of available drill

core from the footwall.

Structures

From the work done to date, the dominant structure within the rock mass is assessed to be

the foliation within the gneiss, schist and amphibolites units forming the hangingwall,

mineralised zone, and footwall respectively. This foliation is parallel to the lithological

boundaries and generally dips at 40° to 45° to the north-west. The geological sections

produced by Amarillo suggest that in places, the lithological boundaries undulate with dips

ranging between 20° and 70°, and it is reasonable to assume that the foliation will do the

same. In the absence of any oriented discontinuity measurements from core logging it is not

possible to confirm this assumption at this stage of the project.

At the boundary between the mineralised zone and the footwall, crenulated foliation and

increased fracturing are observed, in a zone which is typically 1 m thick or less. The nature of

the crenulation suggests that it is shear-induced and marks a sheared contact between the

footwall and mineralised material.

The other significant structure noted to date is the large, planar structure noted by BVP on the

hangingwall of Cava Sul. This structure dips at 65° to 70° to the south-east and has a

persistence of at least 30 m along the hangingwall of Cava Sul (terminations are not seen). If

this structure is one of a family of parallel structures, their presence will need to be taken into

account when deriving design parameters for the hangingwall of the proposed pit.

Groundwater

No information has been provided about likely groundwater conditions or predicted drawdown

during mining. As water currently fills both Cava Norte and Cava Sul to approximately 10m

below current ground level, it is likely that the current water table is at the same level. The

poorly fractured nature of the ground, particularly the hangingwall, means that dewatering

drain holes may need to be drilled to aid lowering of the water table. For the preliminary slope



stability assessment carried out in this study, it will be assumed that adequate dewatering will

be feasible during mining. Given the massive/sparse fracturing of the hangingwall gneiss it

would be a reasonable assumption that the aquifer system would be controlled by the

fractured / sheared contact between the hangingwall and footwall and mineralised zone;

further supported by the depth of weathering differences.

Assessment of likely controlling factors for footwall and hangingwall stability

Footwall

The footwall location and overall slope angle are likely to be controlled by the geometry of the

mineralised zone as this is interpreted to dip at around 40° to the northwest. The foliation also

dips at around this angle. Thus, the controlling factor for the stability of the footwall is

assessed to be the shear strength of the foliation in the amphibolite rock unit.

Hangingwall

The hangingwall rock mass is composed principally of fresh, very strong gneiss with jointing

which is either very widely spaced or massive. The dominant structure (foliation) is interpreted

to dip into the wall at 40° to 45°; the only other significant structure dips towards the southeast

at 65° to 70°. Planar failure along this steeper-dipping structure would not be kinematically

feasible unless the batter angle or inter-ramp slope angle were steeper than this.

Given the orientation of the foliation with respect to the proposed pit wall, it is not

kinematically feasible that shearing along foliation would be a potential failure mechanism on

this wall. Toppling failure could be a possibility, subject to foliation spacing and shear strength

along side and basal joints, and will be considered.

The remaining failure mechanism for consideration is the potential for shear failure through

the rock mass. Given the work done to date, and making assumptions about slope drainage,

a ‘first-pass’ empirical analysis using the Haines-Terbrugge chart should enable a factor of

safety for this mechanism to be assessed.

16.1.5 Slope Stability Assessment

Provisional Inter-ramp slope stability - Hangingwall

Inter-ramp angle – empirical analysis

Coffey Mining has undertaken an assessment of the core logs provided in electronic format by

Amarillo, and, accepting that the rock strength indices should be in the range R4 to R5, as

suggested by BVP, has made a preliminary assessment of the Laubscher Mining Rock Mass

Rating (MRMR) for the rock mass. Adopting the mid-range parameters for number of joint

sets and their condition, and assuming moderate water pressures (all believed to be

conservative assumptions), it is Coffey Mining’s assessment that an MRMR of 60

(representative of a “good rock mass”) could be adopted for the rock mass. Using the Haines

and Terbrugge design chart, this suggests that, for a factor of safety of 1.5, an inter-ramp



slope angle of 55° may be appropriate for preliminary design purposes for the hangingwall.

This inter-ramp slope angle (IRSA) would be appropriate for heights of up to 100m; for the

deeper part of the pit it will be necessary to de-couple the upper and lower parts of the slope

to limit the height to 100 m. It may be possible to achieve this by careful siting of the ramp

within the hangingwall, leading to an overall slope angle of around 45° (incorporating the

ramp), and with a preliminary factor of safety of 1.5. The design chart suggests that these

parameters are at the limit of what may be designed using empirical methods alone; it is

anticipated that the feasibility study geotechnical analysis will enable more rigorous analytical

techniques for slope stability to be employed. The current technique is empirical and uses an

experiential basis for slope design. No account is taken of structure. In the case of the Cava

Sul hangingwall, the ‘discontinuity plane of large persistence’ noted by BVP could invalidate

this approach if its dip is less than the proposed batter angles.

Measurement of this structure indicates that, at surface, its dip is around 65°, but the shape

indicates that it may steepen with depth. If there are multiple discontinuities within the

hangingwall with this orientation, then a batter-berm configuration would have to be adopted

which did not undercut these discontinuities. If the discontinuity is an isolated one, then a

steeper batter berm configuration, with an inter-ramp angle of 55° may be possible. Various

batter / berm configurations, taking account of the discontinuity’s control of batter angle are

given in Table 16.1.5_1 below. Note that Coffey Mining’s minimum recommended berm width

is 8.5 m (10 m berm widths are preferred for batter heights of 30 m), unless good blasting

control which does not compromise berm crests can be demonstrated.

Table 16.1.5_1

Mara Rosa Project Pre-feasibility Study Geotechnical Report

Hangingwall batter-berm configurations for different discontinuity dip angles.

Discontinuity dip angle 65° 65° 70° 67°

Bench Height (m) 20 30 30 30

Berm width (m) 8.5 10.5 8.5 8.5

Batter angle 65° 65° 70° 67°

Inter-ramp Angle 48.3° 50.8° 55° (1) 55° (2)

Notes: (1) Configuration required initially if discontinuity angle remains at 65° until blasting control can be achieved. (2) Required configuration if discontinuity angle remains at 65° with depth – may not be achievable unless very good blasting control can be maintained (3) Preferred configuration, assumes discontinuity angle steepens to 70° with depth (4) Configuration if discontinuity angle steepens to 67° with depth

It can be seen from the above table that the presence or absence of major discontinuities will

be the critical factor in determining whether or not an inter-ramp slope angle of 55° is in fact

achievable. The proposed drill holes for the Geotechnical Feasibility-level Study have been

designed with the priority of determining this in mind.



Toppling failure analysis

Toppling failure on the hangingwall is not considered likely to be a high risk. Preliminary

toppling failure analysis was undertaken using Coffey Mining proprietary software to confirm

that this was the case, and to identify to the variation of which parameters, the model was

most sensitive. The model used a 55° inter-ramp slope angle for a slope height of 200 m, with

the foliation dipping at 40° into the wall. The minimum angle allowed by the software of 50°

was used for the basal plane. The parameters varied were the friction and cohesion of the

basal plane and the side joints and the foliation spacing. For the base case these were set at

200 kPa and 40° for both surfaces and 10 m for the foliation spacing. The base case factor of

safety was modelled to be 1.7. The parameter which had the greatest effect on modelled

factor of safety was shown to be the basal plane cohesion; if this was less than 110 kPa then

a modelled factor of safety of less than 1.3 was returned. It was assessed from these results

that, with the present knowledge of the hangingwall conditions, toppling failure is not likely to

influence the stability of the overall slope.

The most aggressive of the batter-berm configurations quoted in Table 1 (70°, 30 m high

batters) was also checked for toppling failure. A factor of safety of 1.3 or greater is modelled

when the basal plane cohesion was greater than 50 kPa.

Provisional Inter-ramp slope stability - Footwall

BVP’s recommended overall footwall slope angle of 40° to 45° (in fresh rock) would probably

involve undercutting foliation on a batter scale, which, if the rock mass is weak along the

foliation, would lead to numerous batter scale failures; however this may be regarded by

Amarillo as being operationally acceptable if the slope design includes adequate berm width

for rock fall protection. Assuming that a degree of batter scale failure is acceptable, it is

Coffey Mining’s assessment that an inter-ramp slope angle of 40° to 45° may be appropriate

in the fresh rock. Given the weakness imparted by weathering to planes of foliation, it is

Coffey Mining’s recommendation that a batter angle of 40° be adopted for preliminary design

purposes in the weathered part of the footwall. Preliminary batter-berm configurations

suggested for the footwall are given in Table 16.1.5_2 below.

Table 16.1.5_2

Mara Rosa Project Pre-feasibility Study Geotechnical Report

Footwall batter-berm configurations.

Fresh rock Weathered zone

Bench Height (m) 20 20

Berm width (m) 8.5 10

Batter angle 52° 40°

Inter-ramp Angle 40° 31°



The batter-berm configurations for the fresh rock are based on the assumption that any batter

scale failures are likely to be initiated by blast damage, and thus be limited to the upper part of

each batter or berm crest. The strong and relatively weakly-jointed nature of the amphibolites

observed (from limited samples) during Coffey Mining’s recent site visit, suggest that failure of

an entire batter slope along a 40° dipping foliation would not be expected. This will need to be

confirmed by shear strength testing of the amphibolites along foliation in the geotechnical

Feasibility Study program.

The berm width recommended in the table above is based on the assumption that any batter-

scale failures will be limited to the uppermost 15 m of the batter face, for which the catch

capacity is assessed to be adequate. A full-batter failure would be caught completely by two

successive berms.

16.1.6 Pre-feasibility conclusions

Based on the work done by BVP and Coffey Mining’s review and independent assessment of

the data presented to date, Coffey Mining’s conclusions are that:

No further ground investigations should be necessary for a geotechnical study at PFS

level.

Assessment of the geological and geotechnical data presented suggests that BVP’s

preliminary slope design recommendations may be appropriate for a PFS.

The hangingwall gneiss may be characterised as a very good quality rock mass with few

discontinuities and a very shallow depth of weathering.

The properties of the footwall amphibolites and mineralised schists are less certain, but

the controlling factor for stability studies is assessed to most likely be the shear strength

parallel to foliation.

The major structural feature of the deposit is the ductile shear zone which strikes

northeast-southwest and dips between 40° and 55° to the north-west. Few other

discontinuities are observed in field exposures, but a south-eastward dipping structure

with a long persistence was observed by BVP during their site visit. There is some

undulation of the dominant foliation, and the geologic sections of the mineralised zone

provided by Amarillo suggest that sigmoidal shear patterns may well be encountered

within the mineralised zone; possibly extending to the footwall.

The conclusions and recommended pit slope design parameters assume that adequate

slope drainage can be achieved during construction, either by natural drainage of the pit

walls, or by using weep holes to dewater.



16.2 Mining Study

The mine pit design incorporated resource data provided by AEFS into a block model from

which a preliminary pit design was created. The pit design was then optimised using Whittle

Four-X software and specified economic and production input criteria, detailed below. The

resulting pit shell was then incorporated in to Surpac design software to develop the annual

mining plans for the seven-year life of mine.

16.2.1 Pit Optimisation

Block Model and Resources

Block Model Definition

Table 16.2.1_1 shows the Posse Deposit block model definition with block dimensions and

Table 16.2.1_2 shows the model attributes.

Table 16.2.1_1

Mara Rosa Project

Block Model Definition

Direction Minimum Maximum Number of

Blocks Block Size

(m) Minimum Block

Size (m) Coordinate X 695,787.5 695,790.0 65 25 5

Coordinate Y 8,453,787.5 8,453,790.0 57 25 5

Coordinate Z 0.0 470.0 47 10 2

Table 16,2,1_2

Mara Rosa Project

Block Model Attributes

Attribute Name

Type Description

au_g/t Float Gold grade (g/t) modeled by MIK Unit Character Domain : Main,FW,HW and Waste class Character Reserve Classification by Ellipsoid Method class2 Character Regularized Reserve Classification dens_oti Float Specific Gravity (t/m³)

Resource Estimation

The mineral resources statement above a cutoff grade of 0.5 g/t declared for the Posse

Deposit (Section 14) is summarised in Table 16.2.1_3.



Table 16.2.1_3

Mara Rosa Project

Grade and Tonnes Above Specified Cutoff

Cutoff (g/t)

Resource Category

Domain Volume

(m³) Tonnage

(t) Gold (g/t)

Gold (oz)

Gold (%oz)

0.5 MEAS TOTAL 2,001,000 5,463,000 2.04 358,300 26.9

0.5 IND TOTAL 5,638,000 15,393,000 1.65 816,600 61.3

0.5 INF TOTAL 1,330,000 3,630,000 1.34 156,400 11.8

0.5 TOTAL TOTAL 8,969,000 24,485,000 1.69 1,330,400 100.0

Pit Optimization Parameters

Mining Cost

For the purposes of the pit optimization study, mining costs have been defined based on other

similar gold projects and similar active gold mine operations, in this case, based on thecosts

of the São Francisco Mine of Aura Minerals in Mato Grosso State,Brazil (Table 16.2.1_4).

Table 16.2.1_4

Mara Rosa Project

Mine Costs Distribution

Rock type Drilling (US$/t)

Blasting (US$/t)

Loading (US$/t)

Hauling (US$/t) *

Auxiliary Cost

(US$/t)**

Total (US$/t)

Ore 0.4 0.3 0.4 0.6 0.1 1.8

Weathered Waste Rock 0.4 0.2 0.3 0.4 0.1 1.4

Fresh Waste Rock 0.4 0.3 0.4 0.6 0.1 1.8 * Hauling cost does not include a factor for bench depth or haulage distances out of the pit.

** Mine Auxiliary cost includes mine geology, mine planning, mine dewatering and rehabilitation

Slope Angle

Based on preliminary geotechnical studies by BVP Engenharia and Adam Beer of Coffey

Mining, three sets of slope angles for the pit optimization have been defined as shown in

Table 16.2.1_5:

Table 16.2.1_5

Mara Rosa Project

Geotechnical Parameters Summary

ROCK Schist Weathered Rock Gneiss Direction Interval 50º - 230º For all directions 0º - 50º 230º-360º Bench Height Angle 47º 40º 67º 67º Berm Width 5.2 m 7.7 m 5.1 m 5.1 m Inter-ramp Slope Angle 40º 40º 55º 55º Overall Slope Angle 40º 40º 49º 49º



Geometric and Economic Parameters for Pit Optimization

Table 16.2.1_6 shows the parameters used for pit optimization, defined from Amarillo

information or according to the experience of Coffey Mining.

Table 16.2.1_6

Mara Rosa Project

Geometric and Economic Parameters for Pit Optimization

Block Model Au modeled by MIK (Median Indicator Kriging) 2011Base Price ( U$/oz) 1100 Range of Revenue Value (US$) 330 – 2200 step 55

Overall Slope Angle (º) Hard Rock : 49º for HW and 40º for FW * Weathered + soil : 40º for all directions

Density (t/m³) 2.73 for ore and hard rock, 2.40 for weathered and 1.8 for soil Mining Recovery (%) 97 Ore Dilution (%) 3 Mining Cost (US$/t) 1.80 for ore and hard rock and 1.40 for soil and weathered rock Plant Recovery (%) 92.0 Plant Cost (US$/t feed) 10.70 G & A Cost (US$/t feed) 1.54 Sale Cost – 5.09 % of Price( US$/oz) 56 Discount Rate (%) 8.0 Plant Throughput (Mtpa) 2.5 Initial Capital Cost(US$ M) 170.0 Annual Replacement Capital Cost (US$ M) 0.2

Pit Optimization

The results of this exercise produced a family of mathematical pit results which are presented

in Table 16.2.1_7 and in Figure 16.2.1_1.

Pit Option 15 was selected as the optimal pit with revenue factor of 1.0(=US$1,100/oz). This

pit shell contains approximately 19.3 Mt of mineralized material which ensures an adequate

life of mine as well as showing an average NPV of approximately US$210 M.



Table 16.2.1_7

Mara Rosa Project

Mara Rosa Reserves - Results of Optimization in Whittle (Slope :49 for HW / 40 for FW ) - Total Reserves - Model MIK 2011

Pit# Gold Price Oz. Au Mined (koz)

Oz. Au Net Sold

(koz)

Average Grade (g/t)

ROM Cut off

Mill Cutoff Strip Ratio (:1)

Mine Life (a)

Discounted Cash Flow

(US$M

Undiscounted Cash Flow

(US$M)

Waste Mt

Ore Mt Moving

1 330 208 192 2.49 1.33 1.85 1.62 1.0 (25.73) (14.15) 4.22 2.60 6.82 2 385 311 286 2.26 1.14 1.67 2.13 1.7 32.08 52.46 9.11 4.29 13.40 3 440 367 338 2.06 1.00 1.45 2.07 2.2 57.89 84.67 11.49 5.55 17.05 4 495 429 395 1.95 0.89 1.31 2.24 2.7 84.89 118.90 15.32 6.85 22.17 5 550 521 480 1.83 0.80 1.22 2.59 3.6 119.64 165.68 22.97 8.88 31.86 6 605 579 533 1.73 0.73 1.12 2.68 4.2 136.95 190.99 27.88 10.42 38.30 7 660 663 610 1.68 0.67 1.08 3.18 4.9 160.13 225.58 39.01 12.27 51.28 8 715 819 753 1.65 0.62 1.09 4.25 7.1 192.91 284.89 65.48 15.39 80.87 9 770 837 770 1.64 0.57 1.02 4.32 7.3 195.99 290.56 68.47 15.85 84.31 10 825 912 839 1.65 0.53 1.00 4.98 8.5 206.05 312.03 85.75 17.24 102.99 11 880 916 843 1.64 0.50 0.94 4.98 8.6 206.54 313.15 86.57 17.38 103.95 12 935 955 879 1.63 0.47 0.91 5.36 9.2 208.99 319.73 97.51 18.19 115.70 13 990 981 903 1.61 0.44 0.87 5.49 9.5 210.09 322.40 104.00 18.94 122.94 14 1,045 995 916 1.61 0.42 0.83 5.65 9.8 210.45 323.54 108.61 19.24 127.85 15 1,100 997 917 1.61 0.40 0.79 5.65 9.8 210.45 323.59 109.03 19.30 128.3316 1,155 1,014 933 1.60 0.38 0.76 5.85 10.2 210.21 323.42 115.18 19.70 134.88 17 1,210 1,023 941 1.60 0.36 0.74 5.98 10.4 209.90 322.83 119.00 19.90 138.90 18 1,265 1,030 948 1.60 0.35 0.71 6.06 10.6 209.44 321.86 121.73 20.09 141.81 19 1,320 1,031 949 1.59 0.33 0.68 6.07 10.6 209.34 321.65 122.19 20.13 142.32 20 1,375 1,034 951 1.59 0.32 0.65 6.09 10.7 209.08 321.06 123.19 20.22 143.41 21 1,430 1,066 981 1.58 0.31 0.65 6.66 11.7 205.31 312.36 139.67 20.98 160.64 22 1,485 1,078 992 1.57 0.30 0.64 6.84 12.1 203.70 308.34 145.78 21.31 167.08 23 1,540 1,130 1,040 1.58 0.29 0.66 7.97 14.0 196.25 288.89 177.26 22.24 199.49 24 1,595 1,130 1,040 1.58 0.28 0.64 7.97 14.0 196.22 288.78 177.38 22.25 199.63 25 1,650 1,130 1,040 1.58 0.27 0.62 7.98 14.0 196.19 288.70 177.48 22.25 199.72 26 1,705 1,131 1,040 1.58 0.26 0.60 7.98 14.0 196.14 288.56 177.60 22.26 199.86 27 1,760 1,131 1,041 1.58 0.25 0.58 7.99 14.0 195.99 288.11 178.03 22.29 200.32



Figure 16.2.1_1

Mara Rosa Resources - Results of Optimisation in Whittle

208

311

367

429

521

579

663

819 837

912 916955

981 995 997 1.0141.0231.0301.0311.0341.0661.078

1.1301.1301.1301.1311.1311.1481.1491.1541.1541.155

1.155

(50)

-

50

100

150

200

250

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33

PIT

To

nn

es (

Mt)

-50

50

150

250

350

450

550

650

750

850

950

1050

1150

1250

NP

V (

U$x

1000

)

Waste (Mt) Ore (Mt) Cash Flow(U$x1000) NPV (U$x1000) Au (Koz)

Selected Pit



Results of Optimization in Whittle

The selected pit was imported into Surpac software as a surface digital model format file.

The block model was evaluated with Surpac Model Report tools, using the same cutoff grade

as for the resource estimation. Table 16.2.1_8 shows the results of the estimation which has

the effect of elevating Measured and Indicated Mineral Resources to Proven and Probable

Mineral Reserves.

Table 16.2.1_8

Mara Rosa Project

Whittle Pit Tonnage and Grade Summary - Pit US$1,100 / oz - Au > 0.50 g/t

Reserves Category

Domain Volume (m³) Tonnage (t) Au (g/t) Gold (oz)

HW 319,000 869,400 0.71 19.700 Proven MAIN 1,535,100 4,189,900 2.44 328,600 FW 137,100 373,500 0.79 9,400 TOTAL 1,991,200 5,432,800 2.05 357,800 HW 1,341,800 3,658,500 1.15 135,700 Probable MAIN 2,806,700 7,662,100 1.99 489,000 FW 221,500 604,600 0.86 16,700 TOTAL 4,370,000 11,925,200 1.67 641,400 HW 1,660,800 4,528,000 1.07 155,400 Total MAIN 4,341,700 11,851,900 2.15 817,600 Prov+Prob FW 358,600 978,100 0.83 26,100 TOTAL 6,361,100 17,358,000 1.79 999,100

Life Of Mine – Limits Concepts of Selected Pit

Due to the geometry of the mineralized body and existing drill hole spatial distributions, some

Inferred Mineral Resources exist at depth (below pit bottom) which with an additional infill

drilling program may be reclassified.

Consequently, it is recommended that more infill drilling be undertaken to confirm this

hypothesis, so that the development of an underground mine in the future can be considered.

Table 16.2.1_9 shows those Resources internal to Pit which are in addition to the Mineral

Reserves presented in Table 16.2.1_8. These Resources total 5.6 Mt Measured and

Indicated at an average grade of 0.37 g/t and 0.3 Mt Inferred at an average grade of 0.84 g/t.



Table 16.2.1_9

Mara Rosa Project

Waste Summary - Pit US$1,100 / oz - Au < 0.50 g/t

Resources Category

Domain Volume (m³) Tonnage (t) Au (g/t) Gold (oz)

HW 423,100 1,152,700 0.35 13,100 Measured MAIN - - - - FW 208,900 564,100 0.38 6,900 TOTAL 632,000 1,716,800 0.36 20,000 HW 1,101,900 3,000,600 0.37 35,400 Indicated MAIN 29,300 79,900 0.42 1,100 FW 293,700 801,700 0.37 9,600 TOTAL 1,424,800 3,882,200 0.37 46,100 HW 1,525,000 4,153,300 0.36 48,500 Total MAIN 29,300 79,900 0.42 1,100 Meas+Ind FW 502,600 1,365,900 0.38 16,500 TOTAL 2,056,800 5,599,000 0.37 66,100 HW 73,600 200,700 0.52 3,400 Inferred MAIN 31,800 86,400 1.78 4,900 FW 13,900 37,900 0.36 300 TOTAL 119,300 325,000 0.84 8,600

To establish the boundaries of areas for core infrastructure, and in consensus with Amarillo,

the pit with Inferred blocks were considered as “Ore” and used to define the limit of

condemnation.

Also the extension of the ore body to the southwest (mineralized trend) was considered as the

limit of condemnation.

16.2.2 Pit Design

The pit was designed with Surpac Pit design tools adopting the geotechnical and operational

criteria shown in Table 16.2.2_1. Figure 16.2.2_1 shows the proposed ramp dimensions.



Geotechnical and Operational parameters of Pit Design

Table 16.2.2_1

Mara Rosa Project

Geotechnical Parameters of Pit Design

ROCK Gneiss (HW side) Rock) Schist (FW side) )Rock) Weathered RockBench Height (m) 20.0 20.0 10.0 Inter-ramp Slope Angle (º) 55.0 40.0 40.0 Bench Face Angle (º) 67.0 47.0 40.0 Minimum Work Bench 25 x25 25 x25 25 x25 Berm Width (m) 5.1 5.2 7.7 Ramp and Roads Parameters

Ramp Width (m) 20.0 Ramp Width Last Bench (m)

15.0

Ramp Gradient (%) 10.0 Minimum Radius Curvature (m)

30.0

Berm Drainage Gradient (%)

1.0

Figure 16.2.2_1

Ramp Safety and Drainage

Pit Design with Surpac Design Tools

The Pit design was performed using Surpac interactive design tools facilities, contouring the

mathematical pit and ore bodies displayed by bench level. The major challenge was to

design the ramp system with 20 m width roads, in such a way to keep the overall slope angle

in agreement with the Pit optimization and to minimize waste removal.



After considering options, it was decided to design a single ramp system for both north and

south pits as this provided optimal benefits, including:

Less waste rock removal;

Improved reliability in mine operations, avoiding the construction of two ramp systems in

an area of greater slope instability (Footwall side);

The 360 m level of the North Pit allowed other benefits:

A safe area for the parking of mining equipment during blasting activities;

A water sump to collect rain water from the North and South pits and the installation of

one pumping system for mine dewatering and pumping to the water dam.

Figure 16.2.2_2

Designed Pit with Single ramp system to both pits

It is also worth noting that the saddle between the proposed North and South pits is

comprised of mineralised material currently only classified as Inferred Mineral Resources.

Additional evaluation drilling can be anticipated to improve the level of confidence to the

extent that declaration of Indicated or Measured Mineral Resources may be possible with

consequent positive impact on pit design.

Pit design and conciliation

Figure 16.2.2_2 above shows the designed pit, Table 16.2.2_2 shows the run of mine (ROM)

tonnage and grade, and Table 16.2.2_3 shows waste and resource tonnages. The quantities

differ from, and supercede, the optimised pit calculations.



Table 16.2.2_2

Mara Rosa Project

Tonnage and Grade Summary - Pit US$1,100 / oz - Au> 0.50 g/t

Reserves Category

Domain Volume (m³) Tonnage (t) ROM Au (g/t) Gold (oz)

HW 319,000 869,400 0.71 19,700 Proven MAIN 1,508,500 4,117,700 2.42 320,700 FW 138,900 379,300 0.76 9,300 TOTAL 1,966,500 5,366,400 2.03 349,700 HW 1,356,100 3,697,500 1.15 137,200 Probable MAIN 2,725,200 7,439,600 1.97 470,700 FW 224,700 613,300 0.85 16,800 TOTAL 4,306,000 11,750,400 1.65 624,800 HW 1,675,100 4,566,900 1.07 156,900 Total MAIN 4,233,700 11,557,300 2.13 791,400 Prov+Prob FW 363,600 992,600 0.82 26,100 TOTAL 6,272,400 17,116,800 1.77 974,400

Table 16.2.2_3

Mara Rosa Project

Waste and Resources Summary - Pit US$1,100 / oz - Au < 0.50 g/t

Resources Category

Domain Volume

(m³) Tonnage (t) Au (g/t) Gold (oz)

HW 422,900 1,152,200 0.35. 13,100 Measured MAIN - - - - FW 198,700 541,300 0.38 6,600 TOTAL 621,600 1,693,500 0.36 19,700 HW 1,148,000 3,126,700 0.37 37,000 Indicated MAIN 28,600 78,100 0.42 1,100 FW 301,900 824,200 0.37 9,900 TOTAL 1,478,500 4,029,100 0.37 47,900 HW 1,570,900 4,278,900 0.36 50,100 Total MAIN 28,600 78,100 0.42 1,100 Meas+Ind FW 500,600 1,365,500 0.38 16,500 TOTAL 2,100,100 5,722,600 0.37 67,600 HW 88,400 241,000 0.57 4,400 Inferred MAIN 26,500 72,100 1.77 4,100 FW 12,500 34,100 0.35 400 TOTAL 127,400 347,200 0.80 8,900 Waste 48,740,500 131,085,200 0.10 420,300 Total Waste and low grade resources

50,967,900 137,155,000 0.11 496,800

A tally of removed waste and mined ore indicates an overall life-of-mine (LOM) stripping ratio

of 8.01:1.

16.2.3 Mine Scheduling

Mine Scheduling Assumptions

Mine Scheduling was performed to maximize NPV results, provide better productivity of mine

operations, security of operations and guarantee of slope stability.

The following assumptions were adopted



Waste/ore definitions considered in this phase were:

Only blocks classified as Measured and Indicated Resource with grades higher than

0.5 g/t were considered “Ore”;

Blocks with grades lower than 0,5 g/t, blocks classified as Inferred Resources and

waste rock were considered as “Waste”;

An ROM production rate of 2.5 Mtpa of ore to be delivered to the processing plant ;

A ramp-up factor of 75% in the first half and 100% in the second half of the first operating

year was modelled, with a resulting first year scheduled production of 2.19 Mtpa;

Pre-stripping: waste rock removal to achieve six months of exposed ore and the mining

of 200,000 t of ROM to perform blasting tests, ore characterization studies and industrial

plant tests;

A plan to maintain 6 months of exposed ore for the next period;

Prioritize mining in the North pit focusing on rapid depletion of reserves, and in this area

construct a definitive pumping system to dewater both the South and North pits. Also the

level between 320 m to 360 m must be filled with waste material from the South Pit and,

at level 360 m, there must be constructed one escape area for low mobility equipment, to

be used during detonation hours;

Waste deposition to Waste Dump 1 during the initial period to reduce haul distances,

trying to maintained exit ramp near to Waste Dump 1;

Assumed operation parameters listed in Table 16.2.3_1.

Table 16.2.3_1

Mara Rosa Project

Operational Parameters for Mine Scheduling

ROCK Gneiss Schist Weathered Rock Bench Height (m) 10.0 10.0 10.0 Bench Face Angle (º) 70.0 70.0 60.0 Minimum Berm width for each PushBack(m) 10.0 10.0 10.0 Operational Berm Width (m) 40.0 40.0 40.0

Mine Scheduling Phase Design

The mining sequence scheduling was determined by the mining phases and was defined by

time and geometric constraints; these represent surface envelopes for different costs and net

revenue generated from the pit optimization stage.

For the current schedule two mining phases have been considered:



Mining Phase with total depletion of North Pit

Designed Pit.

The adoption of the mining phase with total depletion of the North Pit forces the prioritization

of mining and depletion of the North pit;

The mine scheduling was processed in Mine Sched 6.1 software;

The detailed design includes the generation of crest, toe and ramp lines for each bench in

order to keep the geometry recommended for the final pit, avoiding the mining of waste rock

outside the ultimate pit limits.

Table 16.2.3_2 summarizes the tonnages of each mining period, including gold grades and

contents for in situ and ROM, including 3% dilution by waste rock.

Table 16.2.3_2

Mara Rosa Project

Mine Scheduling

Tonnage Grade in situ Grade ROM

Period Waste kt

Ore kt Strip ratio

Total moved

kt

Au g/t Au kg Au oz Au g/t Au kg Au oz

PS 11,003 200 11,203 1.04 208 6,700 1.01 202 6,500

Year 1 18,139 2,298 7.89 20,438 1.67 3,830 123,100 1.62 3,715 119,400

Year 2 24,324 2,404 10.12 26,729 1.94 4,667 150,000 1.88 4,527 145,500

Year 3 24,579 2,451 10.03 27,030 1.91 4,671 150,200 1.85 4,531 145,700

Year 4 23,660 2,440 9.70 25,958 1.75 4,264 137,100 1.69 4,136 133,000

Year 5 22,961 2,517 9.12 25,365 1.59 3,992 128,300 1.54 3,872 124,500

Year 6 9,790 2,444 4.01 12,241 1.69 4,138 133,100 1.64 4,014 129,100

Year 7 2,733 2,361 1.16 5,173 1.92 4,542 146,000 1.87 4,406 141,600

Total 137,188 17,117 8.01 154,137 1.77 30,309 974,400 1.72 29,399 945,200

The Figure 16.2.3_1 shows the evolution of Waste, Ore, Total Moved, Strip Ratio, grade in

situ for each period.



Figure 16.2.3_1

Mine Scheduling

‐

2

4

6

8

10

12

‐

5,000

10,000

15,000

20,000

25,000

30,000

PS Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7

Grade and Strip ratio

Mass (kt)

Period

Waste Ore Total moved Strip ratio Au ppm

Full details of the design of the eight annual periods are presented in Fonseca and Horta

(2011). Pre-stripping and final year pit geometry are illustrated in Figures 16.2.3_2 and

16.2.3_3.



Figure 16.2.3_2 Pre-Stripping

Figure 16.2.3_3

Period Year 7 - Final Pit Geometry



16.2.4 Waste Rock, Ore and Low Grade Stockpiles

The stockpiles of waste rock, which consist of both altered rock and low-grade ore, will be

built up around the mine pits. All of the material will be transported from the pits to the

unloading areas by truck.

Coffey Mining prepared and conducted the studies that provided the geotechnical parameters

necessary to design the stockpile layout, with later approval by Amarillo Gold with respect to

the locations of the stockpiles and their slope construction.

Volume of Material to be Stockpiled

The total volume of each type of material that will be generated during the mining process are

in accordance with the production program. They are shown in Table 16.2.4_1

Table 16.2.4_1

Mara Rosa Project

Volume of Stockpiled Materials

Material (Mt) (Mm3)

Waste Rock 118.79 60.92

Low Grade 5.71 2.93

Altered 14.46 8.44

TOTAL 138.96 72.29

The figures in Table 16.2.4_1 were taken directly from the mine production program,

assuming a freely settled density of 1.95t/m3 for each type of rock, except for the altered rock,

which density is 1.71 t/m3.

Stockpile Layout and Design Parameters

In accordance with the project design parameters that were developed by Coffey Mining, the

stockpile location was determined so as to minimize transport distances.

The waste rock dumps will be constructed by piling the material upwards in benches that are

10m high, having a face angle that corresponds to the waste rock’s angle of repose, which is

approximately 37º (1.3H : 1V). This angle is within the safety limits required by environmental

agencies, which also require that topsoil and humus be applied to the surface of the dumps in

order to perform hydroseeding. The berms will be 10m wide and have a cross-slope from top

to bottom. From the top bench, the overall angle of the slope, resulting from the geometry of

the individual benches and the angles of the slope face, will be approximately 23º (2.3H : 1V).

See Figure 16.2.4_1.

The altered waste rock dump will be built with 10 m benches and a bench face angle of 32º

(1.6H : 1V). Once the bench face angles have been smoothed by a bulldozer to correspond to

the angle of repose of the material, the berms will be 10 m wide and have a cross slope from



the crest of the lower bench to the toe of the upper bench. The overall slope angle of the

altered waste dump will be approximately 21º (2.6H : 1V) (Figure 16.2.4_2).

The low-grade ore stockpile will be constructed by piling the material upwards in benches that

are 10 m high, having a face angle that corresponds to the angle of repose of this material,

which is approximately 37º (1.3H : 1V), (Figure 16.2.4_1).

All of the vegetation and topsoil will be removed from the area that corresponds to where the

base of the stockpile will be located. This material will be stored close to waste rock dump

WD-02 for future use in revegetating the stockpiles and the pit.

The waste rock dumps will have haulage ramps with slopes not exceeding 10% and widths of

20 m. The ramps will allow access to all of the waste rock dump benches to allow for their

maintenance and hydroseeding.

Figures 16.2.4_1 and 16.2.4_2 depict the slope geometries for the three types of piles.

Each waste rock dump will have sediment basins for preventing fines from running off directly

into the watercourses that flow through the area.

Figure 16.2.4_1

Waste Rock Stockpile and Low-Grade Ore Stockpile Design Parameters



Figure 16.2.4_2

Altered Waste Rock Stockpile Design Parameters

Waste rock dumps will be located outside of the predefined condemnation limits, towards the

southeast area of the pit, beyond the extension of the ore body trend.

The waste rock dumps were placed as close to the areas of mining activity as possible in

order to minimize the costs of material transport.

The waste rock dumps will be located close to the mining pits. One of the waste rock dumps

will be located west of the pit (Waste rock dump WD-01), while the other will be located south

of the pit, downstream from the tailings dam (Waste rock dump WD-03). Waste rock dump

WD-02, for the disposal of altered rock material, will be located east of the pit. Low Grade

Stockpile LGO, for stockpiling low-grade material, will be located close to the crushing circuit

(Low-grade Stockpile – LGO).

Plans provide that the altered rock waste rock dump (WD-02) be placed east of the pit, setting

it apart from the industrial installations and the main office, in order to minimize the quantity of

dust that reaches these facilities. The low-grade ore stockpile will be placed in an area that is

close to the primary crusher to minimize future ore recovery costs.

The location of each waste rock dump is depicted in Figure 16.2.4_3.



Figure 16.2.4_3

Waste Rock Dump Layout



Table 16.2.4_2

Mara Rosa Project

Waste Dump Dimensions

Waste rock dump Initial Platform Final Crest Final Toe Volume Capacity Area

(m) (m) (m) Mm3 K m2

WD-01 390 510 500 29.54 10³ x m2

WD-03 232 390 210 31.38 624.16

Waste rock dump Total 60.92

Altered rock WD-02 410 510 500 8.44

Low Grade Ore LGO 424 450 440 2.93 245 92

The project design details are shown in Table 16.2.4_2.

Waste Rock Removal and Disposal Program

The waste rock removal and disposal program was prepared based on the volumes and

timing of the ore production schedule and took into account the following criteria:

Waste rock dumps will be constructed in an ascending manner in compliance with BVP

Engineering/Adam Beer technical recommendations;

Based on the origin of the material, the closest and/or lowest waste rock disposal area

was given priority in order to minimize not only the total distance travelled but also to

avoid hauling material uphill;

During the pre-production period and the first year of operation, the LGO material will be

used as a foundation for the construction of the low-grade ore stockpile. This low-grade

ore will be processed during the life of the mine.

Tables 16.2.4_3, 16.2.4_4 and 16.2.4_5 summarize, respectively, the waste rock, low-grade

ore and altered waste rock disposal programs.



Table 16.2.4_3

Mara Rosa Project

Program Implementation – Waste Rock Stockpiling

Schedule Material for Dump

Total Mm3

Pre Stripping

Year 1 Year 2 Year 3 Year 4 Year 5 Years6 & 7

WD-01 29.54 2.22 7.01 10.82 9.49 - - -

WD-03 31.38 - - - 1.56 11.94 11.81 6.07

Total 60.92 2.23 7.01 10.82 11.05 11.94 11.81 6.07

Waste rock dump Disposal Schedule (Mm3)

ID Level (m) Total Mm3

Pre Stripping


WD-01

390 0.17 - 0.17 0.00000 - - - -

400 1.16 0.24 0.92 - (0.00) - - -

410 2.26 0.74 1.52 0.00 (0.00) - - -

420 3.59 1.24 2.35 (0.00) 0.00 - - -

430 4.61 - 2.04 2.56 - - - -

440 4.42 - - 4.42 - - - -

450 3.75 - - 3.75 - - - -

460 3.10 - - 0.09 3.00 - - -

470 2.48 - - - 2.48 - - -

480 1.89 - - - 1.89 - - -

490 1.32 - - - 1.32 - - -

500 0.80 - - - 0.80 - - -

Total 29.54 2.22 7.01 10.82 9.49 - - -

WD-03

430 0.25 - - - 0.25 (0.00) - 0.00

440 1.73 - - - 1.31 0.32 0.10 -

450 4.02 - - - - 3.28 0.74 (0.00)

460 5.64 - - - - 3.80 0.86 0.98

470 5.90 - - - - 2.52 2.20 1.18

480 5.18 - - - - 2.02 2.13 1.04

490 4.18 - - - - - 3.29 0.89

500 3.23 - - - - - 2.48 0.75

510 1.24 - - - - - - 1.24

Total 31.38 - - - 1.56 11.94 11.81 6.07



Table 16.2.4_4

Mara Rosa Project

Program Implementation - Low Grade Ore Stockpiling

Schedule Material for Dump

Total Mm3

Pre Stripping


LGO 2.93 0.14 0.43 0.61 0.57 0.38 0.41 0.38

Total 2.94 0.15 0.43 0.61 0.57 0.38 0.41 0.38



Pre Stripping

Year 01 Year 02 Year 03 Year 04 Year 05 Years 06 and

07

LGO

420 0.35 0.14 0.21 0.00 0.00 0.00 0.00 0.00

430 1.30 0.00 0.22 0.61 0.46 0.00 0.00 0.00

440 1.29 0.00 0.00 0.00 0.11 0.38 0.41 0.38

Total 2.93 0.14 0.43 0.61 0.57 0.38 0.41 0.38

Table 16.2.4_5

Mara Rosa Project

Program Implementation – Altered Rock Stockpiling

Schedule Material for Waste rock dump

Total Mm3

Pre Stripping

Year 1 Year 2 Year 3 Year 4 Year 5 Years &

6 & 7

WD-02 8.44 3.78 2.45 1.06 1.14 - - -

Total 8.44 3.79 2.45 1.06 1.14 - - -



Pre Stripping


WD-02

410 0.03 0.03 - 0.00 - - - -

420 0.30 0.30 - 0.00 - - - -

430 1.29 1.29 - 0.00 - - - -

440 1.84 1.84 - 0.00 - - - -

450 1.55 0.33 1.23 0.00 - - - -

460 1.22 - 1.22 0.00 - - - -

470 0.92 - - 0.92 - - - -

480 0.65 - - 0.14 0.51 - - -

490 0.42 - - 0.00 0.42 - - -

500 0.22 - - 0.00 0.22 - - -

Total 8.44 3.78 2.45 1.06 1.14 - - -

16.2.5 Tailings Storage Facility

The design for the Tailings Storage Facility (TSF) and related Water Storage Facility (WSF)

for the Mara Rosa Gold Project (Saunders, 2011) has been aimed at:



Optimising tailings storage capacity by maximising tailings density; and

Reducing environmental and societal impact.

Whilst the concept of the WSF is intended to achieve:

maximise collection and reuse of decant water; whilst

minimising environmental impacts;

reducing evaporation losses.

Based on the site selection study that was carried out by Coffey Mining, (Coffey, June 2011),

Site 1 has been used as the preferred area in which to develop the prefeasibility study design

(Figure 16.2.5_1). To the east of the TSF basin, there is a stream flowing north on which are

two potential water storage dam sites, one of which could provide storage capacity for the

decant water and stormwater collected from the TSF.



Figure 16.2.5_1

TSF Options Study – Site Layout



Site and topographic information has been provided by Amarillo , together with a hydrological

report, a process mass balance and a preliminary water balance. This information is the

basis for the design work discussed here and has been accepted and incorporated into the

design process without specific verification.

TSF Design Concept

Design Parameters

The design for the TSF is based on the following parameters:

tailings slurry density nominally 59% solids (by weight);

deposited tailings dry density has been assumed to be 1.25 t/m³; and

an estimated beach slope of 1% has been used.

The proposed TSF has been designed in general accordance with the ANCOLD guidelines.

Storage Capacity

The capacity of the TSF is based on the following parameters in Table 16.2.5_1 below.

Table 16.2.5_1

TSF Design Criteria

Total Tailings Production (Mt) 20

Storage Facility Design Life (years) 8

Slurry Density (solids by weight) 59%

No physical or geochemical tailings test work, site drilling or test pitting has been undertaken.

Hence, tailings storage design work has been performed based on assumed parameters.

It is intended that the TSF be constructed in stages over the life of the mine, at two yearly

intervals. The storage capacity for each stage is summarised in Table 16.2.5_2.

Table 16.2.5_2

Storage Capacity and Raise Implementation

Stage No. Dry Mass Stored

(t) Stage Implementation

(Years after start)

1 4,073,000 Start up

2 3,877,000 2.04*

3 3,969,000 4.00

4 4,274,000 6.00

Total 16,193,000

* the construction of Stage 2 may be delayed if the initial mine production is less than the 250,000tpa.



The construction of each stage is determined by the rising beach of the tailings and will need

to be implemented well before the freeboard is less than the design of 2 m. The actual timing

of these would be verified as the project develops.

Hazard Rating and Risk

The Australian National Committee On Large Dams (ANCOLD) has produced guidelines on

tailings dam design defining the hazard rating for tailings storages and develops this into a

Hazard Category that defines the spillway and freeboard requirements for the TSF. The

hazard rating of the Mara Rosa TSF is Significant, particularly with the mine development

being downstream of the TSF; however, this takes into consideration the positioning of the

waste rock stockpile between the main embankment and the mine. This rating has been

assessed based on the following:

No loss of life expected following a TSF embankment failure but the possibility recognised;

Appreciable economic loss recognised (ie significant loss to mine production) following a

TSF embankment failure

To address the risk that is created by the TSF the freeboard requirements are defined as:

storage, above the normal year high pond level, for the 1:1000 AEP storm plus 0.3 m

freeboard; or

worst wet season on record; less water returned to plant plus wave run-up plus 0.3 m

freeboard.

In view of the limited amount of rainfall data available from which to extrapolate the

1:1,000 AEP storm event, the latter criterion was used for the PFS design.

Embankment Design

The preliminary design of the final stage of the TSF perimeter embankment is shown on

Figure 16.2.5_2, along with the waste rock stockpiles (WRS), decant towers and access

causeways. The volumes of material involved in the embankment construction have been

estimated assuming an embankment crest of 10m and batter slopes of 1(v) : 2(h) upstream

and 1(v) : 2.5(h) downstream, although the earlier stages of the embankment have been

designed with a steeper downstream face slope of 1(v) : 1.5(h), should rock for the

construction prove to be scarce. The embankment is designed so that suitable waste rock

can be continuously placed on the downstream face in preparation for the construction of the

raising of the upstream core, once the foundation area has been prepared.



Figure 16.2.5_2

TSF Option 1 General Arrangement



Using the design annual deposition rate of 2.5 Mtpa and an anticipated dry density of the

tailings of 1.25 t/m³, the perimeter embankment crest levels were established from the storage

capacity assessments.

Table 16.2.5_3

Embankment Statistics by Stage

Stage No. Embankment Crest

RL (m) Embankment Total *

Fill Volume (m³)

1 465.0 447,000

2 470.5 337,000

3 474.5 359,000

4 478.0 522,000

Total 1,665,000

*Excludes volumes for decant causeway.

The crest levels in Table 16.2.5_3 include an allowance of 2 m ‘freeboard’ above the

deposited beach of tailings but do not take into account:

any water balance calculations; or

the beach gradient that will provide additional freeboard.

The stability of this perimeter embankment is achieved by the construction of an extensive

downstream portion of Zone 3 material that comprises waste rock from the mining excavation.

The granular material with rock between 300 mm and 600 mm in size will have shear strength

in excess of that of normal soil material, particularly as it is intended to be compacted in 1 m

thick layers by a large bulldozer. This free draining material will also provide support to the

deposited tailings in the situation of an earthquake, at which time the saturated tailings may

tend to liquefy.

The broad, 10 m wide, crest of the proposed embankment that enables safe access by large

mining haul vehicles, contributes to the stability of the geotechnical structure.

Although the downstream portion of the embankment is constructed from free draining rock,

the upstream face forms a low permeability zone of clay material that extends below the

existing ground level. The cutoff trench situated at the upstream toe of the embankment is to

extend to a relatively impervious horizon within the foundations and hence will reduce the

seepage through the upper layers of the natural ground.

The clay, which is to be obtained from carefully selected borrow areas, is to be moisture

conditioned and compacted in thin layers to achieve the lowest permeability practical, both

below ground, in the cutoff trench, and in the upstream face of the embankment. In this way

water emanating from the tailings will be contained within the basin of the TSF and collected

by the underdrainage system that is connected to the base of the decant towers.



Water Management

Efficient reuse of the process water will benefit the project by reducing the requirement for

makeup water that can only be obtained from an aquifer close to the town of Mara Rosa

(HIDROVIA, 2011) or from abundant surface water.

To achieve the maximum recovery, the design of the TSF incorporates a decant system to

pick up the supernatant water from the tailings deposited within the basin. This decant

system will also recover stormwater run-off from the catchment of the TSF and return a

significant portion of this water to the WSF.

The catchment of the TSF extends towards the Mara Rosa Road in the south-west and is

bounded on the northwest and southeast by a range of hills, giving a total catchment area of

130ha. Run-off from the tailings beach will be a significant portion of the rainfall as this is a

relatively impervious surface of fine material and any rain falling on the decant pond will be

recovered to the WSF. The area of the beach will progressively increase with ongoing

deposition, whilst the area of the decant pond will depend on the capability of the water

recovery system. All of these factors need to be taken into consideration in the development

of the water balance for the TSF and WSF.

The conceptual layout of the decants’ causeways is shown on Figure 16.2.5_2 illustrating that

the pond is expected to migrate progressively south-west as the facility is developed. Each

tower is connected to a system of underdrainage that will enable pore water to drain from the

tailings into the base of the tower for recovery as return water. This underdrainage not only

serves to recover water but will facilitate the consolidation of the tailings and increase the final

density of the deposited material.

Although the causeways are intended to be developed in stages, the foundations of the

Stage 2 and 3 towers need to be established during Stage 1. Whilst the initial tower will be

constructed to full height at start-up, the Stage 2 and 3 towers will be raised as the level of the

tailings beach approaches the top ring so that tailings does not enter the pump sump.

Installed in each of the decant towers will be a submersible return water pump connected by

flexible rubber hose to the pipelines leading along the causeways to the perimeter

embankment. These pipelines will deliver the return water to the WFS from where it can be

discharged into a pipeline to the plant process water tank, hence completing the cycle. The

decant tower will also collect runoff from within the TSF basin and this will join the recovered

process water in the WFS.

The capacity of this return water system will be determined in the course of the water balance

calculations, to optimise the water recovered and minimise evaporation and seepage losses.



Because this TSF represents a significant hazard, the design flood will be based on the

highest rainfall event encountered. Rainfall figures were extracted from the hydrological

report.

The highest monthly total rainfall recorded for the TSF location was 835.3 mm in January 1985.

The ‘extreme’ figures will be utilised in the detailed design water balance to determine:

the required flood capacity of the TSF;

optimisation of decant pump and pipeline; and

design capacity of the emergency spillway.

The minimum rainfall figures, which represent an exceedingly dry year, will be used to

investigate the need for additional raw water supply over and above that which can be

obtained from the pit dewatering.

Appurtenant Works

A certain amount of infrastructure is required to successfully construct and operate the TSF

and the decant towers. The TSF design concept also includes the following items which have

not generally been included in the schedule of quantities:

Access roads to the TSF from:

the plant for light vehicles, and

mine pit for waste rock haul trucks.

Tailings pipelines for delivery, within bunded corridor and spill catch-pits;

Slurry distribution pipes and spigots along the perimeter embankment;

Return water pipeline to convey the process water accumulated in the WSF back to the

plant;

Monitoring equipment (included in the schedule) for:

groundwater (standing water level and contaminants); and

embankment stability and settlement.

Facilities required for the operation of the TSF that will be designed by others will be:

Power supply, distribution and switching; and

Instrumentation and pump control for both the tailings and decant water.

As the project moves into the detailed design stage these aspects will be addressed and

included in the construction quantities.



Rehabilitation and Closure

Essentially the rehabilitation and closure of the TSF is intended to achieve, in perpetuity, a

stable structure that the limits the environmental impact of the contained materials whilst

being resistant to erosion and compatible with the surrounding landscape.

The geochemical nature of the tailings and the waste rock will significantly influence the closure

proposals, and these materials have not been tested for their potential to generate acid, in

conjunction with oxygen and water. In order to define the quantities for the conceptual closure

design it has been assumed that it will involve rock fill being placed over the beach of the TSF

and then covered by compacted over burden from the pit; to achieve a convex self draining

profile. Over this surface will be placed a layer of top soil on which natural vegetation be planted.

The post-mining use of the land should be taken in to consideration in the closure plan and

the relevant stakeholders should be consulted early in the process. Their recommendations

for rehabilitation of this structure should be researched and reviewed periodically during the

life of the project and this ongoing process will need to be under the direction of personnel

from the environmental team. A rehabilitation / closure plan will be prepared prior to

decommissioning of the TSF, although the outline proposal for this will be discussed in

subsequent stages of the design process.

Construction Quantities

Table 16.2.5_4 below provides a summary of the embankment construction quantities for the

four stages, with Stage 1 being carried out during the start-up period. The construction of the

decant causeways is to commence at the start up of the construction, with the Stage 1

embankment to be completed to its full height, whilst development of the Stage 2 and 3

causeways will be undertaken over a period of time after deposition has commenced.

Table 16.2.5_4


Construction Materials Quantity

Item Description Units Stage 1 Stage 2 Stage 3 Stage 4

1 Site preparation: clearing vegetation and stripping topsoil

ha 61 30 29 24

2 Excavation of cutoff trench and underdrainage

m³ 37,000 13,000 3,000 3,000

3 Main embankment: Zone 1 fill m³ 128,000 57,000 49,000 62,000

4 Main embankment: Zone 2 filter m³ 8,800 5,700 5,600 4,600

5 Main embankment: Bidim geotextile m² 53,000 37,000 32,000 45,000

6 Main embankment: Zone 3 rock fill m³ 310,000 280,000 310,000 460,000

7 Decant Causeway: Zone 3 rock fill m³ 341,000* 179,000 104,000 48,000

8 Megaflo underdrains m 2,400* 0 0 0

Notes: *Stage 1, 2 & 3 combined Zone 1 fill consists of compacted clay. Zone 2 fill consists of +20mm to -75mm crushed stone or clean natural gravel. Zone 3 fill consists of compacted pit overburden or waste rock.



The timely provision of adequate quantities of the fill materials is dependent on the pit

development and the scheduling of the excavation of the over burden, which will form the

primary constituent of the Zone 3 rock fill.

For completeness, the areas to be rehabilitated for the closure of the TSF have been

estimated along with the volume of materials required to carry out this work.

Water Storage Facility

Although the site selection study was based on storage of process and storm water on the

TSF, this practice leads to the following disadvantages:

excessive embankment height to provide water storage capacity;

provision of additional underdrainage capacity to intercept seepage;

increased evaporation loss of supernatant water due to the large pond size; and

reduction in settled density and shear strength of deposited tailings because:

evaporative drying on the beach is reduced; and

high moisture content of deposited tailings due to less consolidation.

The reduction in shear strength and high moisture content increases the potential for the

tailings to liquefy during an earthquake, reducing the security of the TSF and placing

additional reliance on the embankment to support the contained material.

Hence, the region around the selected site was examined for the potential to develop a water

storage facility (WSF) in close proximity to the TSF were the supernatant water could be

stored prior to use in the plant.

Alternative Sites

To the east of the proposed TSF perimeter embankment is a valley that extends almost

parallel to the embankment toe and within this is an existing water storage dam used for

irrigation. The potential to incorporate the existing dam embankment into a larger structure

was the first alternative to be examined and gave a storage capacity of 351,000 m³ behind a

12 m high embankment.

Further downstream the valley narrows and a further conceptual design determined that an

11 m high embankment would provide storage of 514,000 m³. This downstream location has

a saddle spillway site that would allow for the full supply level (FSL) to be adjusted according

to the required storage.

As the initial estimate has identified a potential shortfall of some 64 m³/h (even including an

allowance for water from pit dewatering) during a dry year, a WSF with the capacity in the

region of 560,000 m³ could provide this water requirement. Whilst this calculation has been



used to provide this volume, the derivation need to be reviewed in the light of the progressing

design, particularly the evaporation from the TSF decant pond and the surface of the WSF.

Since the project requires the provision of a significant amount of make-up water in a dry

year, the downstream site is considered the most appropriate to develop, and the capacity

would be finalised in conjunction with the water balance to be undertaken with the detail

design.

Capacity and Surface Area

The high flood level of the WSF is constrained by the toe of the TSF embankment at its final

height (Stage 4) to a maximum RL of 460 m, which restricts the FSL to RL 459 m. This

coincides with the crest of the hill on the north eastern side of the valley where it is proposed

that the spillway be situated. As the spillway takes advantage of the saddle within the eastern

ridge, the full supply level can efficiently be varied between RL 452 m and RL 459 m which

provides for a range of storage capacity from 22,000 m³ to 514,000 m³ to accommodate the

required volume. Within this range of storage volumes the surface area varies from

18,000 m² to a maximum of 123,000 m². This WSF has a natural catchment area that will

contribute a portion of the stored water.

Embankment Design

It is proposed that the WSF embankments are zoned with a central clay (Zone 1) core and

upstream and downstream faces comprising gravelly clay (Zone 1A), with both these

materials being compacted to 95% of standard maximum dry density (SMDD). The core of

the embankment would extend into the foundations in a cutoff trench that would extend below

original ground level to a relatively impervious horizon of weathered rock or dense material.

The main embankment of the WSF will be approximately 360 m long and contain some

53,000 m³ of Zone 1 clay within the core and cutoff, with a further 45,000 m³ of Zone 1A

material forming the upstream and downstream faces. A feature of this embankment is the

rock fill toe, constructed predominantly of Zone 3 material, on the downstream face that

provides an 8m wide berm at RL 452.5 m and incorporates a filter system to accommodate

seepage, improve stability of the structure and reduce the potential for piping failure through

the embankment. The crest of the embankment is at RL 460.0 m and has a narrow width of

6m, as this is not intended to be a road access. This crest will have a cross fall of 2% to

enable this surface to drain to the basin and will be connected to the intake tower by a precast

concrete walkway.

Whilst the saddle embankment has the same crest level of RL 460 m it is only 200 m long

with a maximum height of 8 m. This embankment shares the ridge with the alignment of the

mine access road and hence, it would be necessary to accommodate this along the crest of

the embankment. It is intended that this be achieved by providing a 10 m wide crest covered

with a wearing course of 300 mm of gravel. It will be necessary to provide edge barriers to

this embankment to mitigate the risk of vehicles driving off the crest.



The design provides for the spillway to be constructed within this embankment, to take

advantage of the topography and discharge down an alternative valley, away from the mine

pit.

Spillway Design

Although it is not intended that the WSF discharge flood water, it is necessary for this dam to

have a spillway so that the embankment is not breached during exceptional conditions. It is

intended that the spillway be provided with a concrete invert at RL 459 m and flow through

this would be guided down a return channel to the stream valley below. To reduce the

potential for erosion of this channel, if it is not excavated into hard material, stone pitching will

be placed over the inverted and the cut faces.

By using box culverts for the structure, traffic on the access road will be able to cross the

spillway even if it is discharging. As the emergency spillway of the TSF will discharge into the

same catchment, the WSF spillway will have the capacity to deal with both flood flows and will

maximise the flood retention by having at least 1m of freeboard.

Outlet Works

Water stored in the WSF will be released into a pipeline leading to the plant for reuse in the

process. Flow in this pipeline will be controlled by a gate valve situated in a concrete

structure built in to the rock fill toe of the main embankment. An alternative gate valve will

control the flow of water from the WSF to the plant. A further butterfly valve positioned

upstream of the other valves will enable repairs to the control valve.

A 600 mm diameter steel pipe is intended to connect the upstream intake tower to the valve-

house at the downstream toe and this flanged pipe will be encased in reinforced concrete.

Within the intake tower will be provision for emergency closure of the outlet pipe in the

unlikely event of failure or need for maintenance of the butterfly valve.


Details of the quantities of the materials required in the WSF construction are contained in

Saunders (2011), however the PFS design does not allow for inclusion of quantities for the

rock pitching or concrete works.

Operation

Tailings will be pumped from the thickener at the plant by way of an overland pipeline that will

be within a bunded corridor that should also have spill catch-pits at regular intervals. At the

TSF perimeter embankment, the delivery line will feed into the distribution main running along

the upstream edge of the crest and this pipeline will be fitted with tees at specific spacing

which can serve as spigots to discharge the slurry, as discussed below.

Water recovery from the facility will be from the decant structures and flow into the decant

tower will be pumped along the decant access way to the perimeter embankment, where it will



join the channel or pipeline leading to the WSF. The process water stored in the WSF will be

released through the outlet works, along a return water pipeline to the plant, at a rate

comparable with the process consumption. Only in exceptional circumstances of high rainfall

will water be discharged from the WSF through the spillway.

Every two years the perimeter embankment will need to be raised by the downstream

method, and this construction which may take a considerable period to complete and should

commence well in advance of the storage capacity being required. The construction of the

upstream core of the embankment could disrupt the distribution of the tailings as this pipeline

will have to be moved, firstly onto the tailings beach and then onto the new embankment crest

once the construction has been completed. These operations could be carried out during a

mill shut down, when the plant is not producing tailings.

Deposition of Tailings

Tailings in the form of a thickened slurry with 59% solids will be discharged sub-aerially onto a

relatively dry beach from multiple single point discharges (spigots) located along the main

embankment crest. In order to achieve the proposed design capacity of the TSF, the

discharge of the tailings slurry should be changed cyclically between the spigots, as

appropriate, to provide uniform development of the tailings beaches towards the basin of the

TSF and position the water pond adjacent to the decant structures.

Depending on the capacity of the tailings pumps, a number of spigots should be operated

simultaneously so that a low flow velocity is achieved. It should be noted that the distribution

pipeline could become silted up if too many spigots are open at the same time. This low

discharge velocity will reduce the potential for the slurry to erode the embankment face,

however; it is desirable that each spigot be fitted with a length of a flexible pipe, so that the

flow can be conveyed down the face to the existing beach.

Water Management

An objective of the TSF management is to reduce evaporation losses and maximise water

return and this can be achieved if the pond is maintained as small as practical, without the

supernatant water becoming turbid. Although wave action will increase the amount of

suspended solids, the depth of the pond will influence the clarity of the recovered liquor. Careful

observation and management is required to achieve an acceptable quality of water. The TSF

could contain a considerable body of water following a significant rainfall event and as much of

this is possible should be transferred to the WSF in as short a time as possible. The minimum

operational freeboard at the TSF perimeter embankment should be maintained at 0.3 m.

The aspects influencing water management are shown on Figure 16.2.5_3 for the TSF and

Figure 16.2.5_4 in respect to the WSF. In both diagrams there is evaporation from the stored

water surface from within the basin; however, these losses can be reduced to a minimum by

storing as much water as possible within the WSF. This is because the depth of storage in



the dam is considerably greater than the pond on the TSF and hence the surface area is less;

proportionately reducing the evaporation.

Figure 16.2.5_3

TSF Water Balance

Figure 16.2.5_4

WSF Water Balance



A number of parameters within the water balance calculations will vary as the project

develops. For example, an area of some 57 ha of the catchment of the TSF will be stripped of

vegetation and topsoil for Stage 1 and the run-off coefficient for this condition will be 0.85,

whilst the remainder of the catchment, with natural ground and vegetation intact, will have a

coefficient of 0.6. However, as this stripped area is covered by tailings the flow will

significantly increase as the relatively impervious surface of fine material of the tailings beach

will cause 80% of rainfall to run-off. Any rain falling on the decant pond will be recovered to

the WSF. The area of the beach will progressively increase with ongoing deposition, whilst the

area of the decant pond will depend on the capability of the water recovery system and the

season.

A detailed theoretical water balance will be conducted for the design report, however;

measurement of the water being sent to the TSF and recovered will serve to establish an

actual water balance for the site from which the behaviour of the two storages can be

assessed. As the data is accumulated prediction of storage volumes could be carried out to

support long-term water management policies for the mine.

Monitoring

Effective management of the TSF depends upon appropriate levels of monitoring and the

implementation of recommended action dependent on those results. Whilst most of the

monitoring is visual, the less frequent exercises listed below involve measurement of

parameters and will contribute to the successful operation of this geotechnical structure.

Frequent inspections should be made of the tailings and water return pipelines, discharge

points, decant system, the position of the supernatant ponds in relation to the water recovery

system and the operation of the underdrainage. The embankments should be inspected once

per day during the day shift whilst the remaining infrastructure should be inspected during

every shift.

Only by regular inspection and appropriate remedial action can the performance of the

water return system be optimised and operational problems avoided.

The requirements for routine inspections should be set out in an Operations Manual (to be

compiled as part of the detailed design phase) and is to be undertaken during each

production shift by an operator or shift supervisor. The inspections should cover the following

aspects of the operation and the observations recorded on a check list:

Tailings slurry pumps (Thickener underflow pumps).

The tailings delivery line and water return pipelines.

Tailings deposition.



Location and size of the water pond.

The decant tower and pump.

The general integrity of the embankment i.e. any cracking or erosion.

Any changes to existing seepage.

The process plant management has the responsibility for verifying that the inspections as

outlined in the Operations Manual have been carried out and monthly inspections of the

tailings storages should to be carried out by process plant management.

Operation, safety and environmental aspects should be periodically reviewed during an

inspection by a suitably experienced and qualified engineer. This inspection should be done

at least every year and enable a periodic review of the scheduling of the next raising.

A number of aspects of the behaviour of the TSF should be monitored on a regular basis and

the following information should be recorded at a minimum on a monthly basis or more

frequently if possible:

Ore treatment, measured in dry tonnes (weekly).

Tailings produced and delivered to the TSF.

Tailings slurry density, measured in percentage solids or slurry water volume (weekly).

Water return from all sources from the tailings storage to the process plant, measured in

cubic metres or tonnes.

Water level in the WSF (weekly).

Recording of climatic data for the site (i.e. rainfall and evaporation).

Environmental monitoring of groundwater bores:

Standing water depths in the bores should be recorded on a monthly basis; and

Water samples, taken from each bore and the WSF, tested for the presence of

particular contaminants (i.e. for water quality) as a minimum on a quarterly basis.

Detailed level surveys should to be carried out at least on an six monthly basis on:

the perimeter embankment settlement monuments; and

tailings beach, as far as this is practical.

Updating of as built survey plans following embankment construction or annually if rock

fill is placed progressively.

This will enable the storage volume that has been used to be reconciled with the tailings

tonnage deposited into the storages to establish an in-situ density of the tailings from



comparison with the assumed design density. The figures, in conjunction with the climatic

data also recorded, will enable the development of a site water balance.

Baseline records should be established for the groundwater depths and water quality prior to

commencing deposition of tailings. Therefore, it would be preferable that the monitoring

bores be implemented during the site investigation and the standing water levels and water

quality measured before mining begins. Additional monitoring should be undertaken after an

exceptional event such as a flood or significant earthquake.

Environmental Risks

Although the design of the TSF, and the incorporation of the WSF, are intended to mitigate

the environmental risks from the tailings storage as far as is reasonably possible, there are

residual hazards that occur at specific times during the life of the TSF. The following sections

discuss these with a view to implementing measures to mitigate the impacts during project

implementation.

Construction

During the start-up and each stage of the construction of the TSF there will be a significant

number of vehicles and plant operating on the site and these will create an impact on the

environment.

Haul and inspection vehicles could generate dust from during a construction period, however;

this can be reduced by the application of the suppressant (at least water) on haul and access

roads and implementing speed limits. The use of appropriate wearing course gravel with a

limited clay content would benefit the project in terms of safety as well.

Inefficient or old engines on trucks and plant produce excessive exhaust fumes and frequently

lose oil but these adverse impacts can be mitigated by using modern and well maintained

plant. Requiring contractors employed on the site to do the same would assist in reducing

environmental damage.

Whilst it is inevitable that a certain amount noise will be generated by the trucks and

construction plant this can be reduced by insisting upon appropriate silencers and exhaust

systems being fitted. The impact on fauna, particularly nocturnal creatures, can further be

reduced by conducting the construction operations only during daylight hours.

In the course of clearing the basin of the TSF there will inevitably be destruction of natural

vegetation and this will equally apply to the haul roads. The extent of the removal of

vegetation can be restricted by demarcating the extent of the construction area prior to the

work commencing. Similarly, a system of permanent haul roads that will service the entire life

of the TSF should be designed in conjunction with the mine pit and these roads set out by

surveyors before the development is undertaken.

Operation



The embankment design discusses the mitigation of seepage from the TSF and this should

form part of the water management strategy for the mine. Should seepage give rise to an

elevated water table downstream of the main embankment, and the water quality of the

groundwater is such that it could damage plant life, measures should be put in place to

reduce this elevated water table. This would predominantly comprise recovery bores that

could pump out the seepage and lower the water table, whilst seepage collection trenches

parallel to the embankment toe serve to intercept shallow flow.

There are potentially two sources of dust that could impact on the site during normal operation

and these are:

dust from the inspection vehicles; and

windblown silt and sand from the dry TSF beach.

The enforcement of speed limits on the access roads to the TSF will reduce the amount of

dust produced by the inspection vehicles, whilst well constructed roads with selected wearing

course material would assist in this.

It is proposed that the slurry be deposited from a sequence of spigots in a cyclic manner and,

by changing the area of beach on which deposition is taking place, the maximum amount of

the TSF surface will be covered by moist or wet tailings. This moisture in the tailings will

reduce the tendency for windblown dust to be generated.

There are two occasions when the uncontrolled release of tailings into the environment can

occur and these are when there is an embankment failure or from a burst pipeline. Whilst

both of these events are unlikely if adequate design has been undertaken, there are ways in

which to reduce the risk of occurrence or mitigate the impact. In Section 5 there is discussion

of the pipeline corridor with containment bunds and catch-pits and this is an accepted manner

in which to confine the contents of a burst or even a leak. This containment system should be

supplemented by a pressure drop cut-out switch for the underflow pump of the thickener.

In contrast, no such simple constraint will contain an embankment failure; however an

adequate design will lower the risk of such an occurrence, but this must be complemented by

appropriate operation of the TSF and monitoring of the embankment behaviour through

inspection and equipment. The design concept includes level monitoring beacons along the

crest and should there be a deflection apparent on these, more advanced equipment can be

installed to measure any material movement.

Process water recovered from the TSF may well contain contaminants and a spill from the

return water pipeline could be detrimental to the environment along the pipeline route and a

similar arrangement can be applied to the return water pipeline (if it is not within the same

corridor). The flow in this pipeline is under gravity and detection of a burst pipeline is

therefore a little more complex requiring a ‘no’ flow detector that is set when the valve is open.



Post Mining

Prior to rehabilitation of the TSF's surface a considerable period of drying out is likely to be

required and during this time the beach will become desiccated and potential for windblown

dust generation considerably increases. Since moisture is no longer being sent to the TSF,

an alternative must be sought to constrain the fine particles and this can be achieved by

applying one of a number of proprietary products specifically designed for this purpose.

Currently it is not known if the tailings or the waste rock is PAF and whether or not there is the

unlikely potential for acid mine drainage to occur. As the design process progresses this will be

investigated, however, the design of rehabilitation and capping of the TSF will take this into

consideration so as to reduce the risk of decreasing pH with the related potential metal ion

migration.

With regular testing of the groundwater seepage would have been detected during the

operation of the TSF and, should it be necessary, a management program can be developed

to recover this water, improve upon its quality and return it to the river system or groundwater.

Such an operation would have to be designed to be sustainable for many years, to take into

account the period over which rehabilitation would be required.

16.2.6 Mine Production and Operating Parameters

Equipment Selection Criteria

The open-pit mining operation at the Mara Rosa Project will utilize the following unit

operations: drilling, blasting, loading and hauling. Ancillary equipment will be necessary to

maintain the operational areas, roads and waste rock piles.

The majority of the mining will be conducted on 10 m benches, with the exception of 30% of

the ore that will be mined on 5 m benches in order to reduce the dilution at the contact areas

between the mineralized zones. Consequently, the drilling and loading equipment were

selected to combine the virtues of high productivity and low unit cost with those of enhanced

mobility and versatility.

Diesel hydraulic excavators, having bucket capacities of 10 m3, will be used in loading waste

rock onto 100 t capacity off-highway trucks, which constitute the main equipment responsible

for the removal of this material. Smaller mobile units include hydraulic drill rigs, for drilling 4"

to 5" holes, and diesel hydraulic excavators with bucket capacities of six cubic meters.

The 4" hydraulic down-the-hole-hammer drill rigs have reverse circulation capability, enabling

them to perform grade control sampling. The 6 m3 capacity hydraulic excavators will be used

mainly in mining the ore, which requires selectivity and mobility. The 5" hydraulic top hammer

drill rigs and 10 m3 capacity excavators will be employed for the removal of waste rock.

Diesel-powered equipment (hydraulic drill rigs and excavators) will exclusively be used in the

mining operations at Mina Posse.



Two 5 m3 capacity front loaders will support smaller operational units that operate in cramped

spaces, open new access roads and remove debris from operational areas, among other

services.

One hundred tonne and fifty tonne capacity trucks were selected, respectively, for

transporting rock waste and ore from the pit. Trucks having these specifications work well with

the hydraulic excavators that were selected, making a total of only five to six loading passes

necessary.

The ancillary equipment is composed of two 464 Hp track-type tractors (similar to a Caterpillar

D9 model). This equipment is used to maintain the pit's operational fronts and waste rock

piles clear of debris, and to clear debris that has accumulated on the berms. A fuel/lube truck,

a motor grader (14’ width blade) and a 20 m3 water truck complete the list of ancillary

equipment.

Amarillo Gold personnel and equipment will conduct the mining and pre-stripping operations.

16.2.7 Operations Timetable

The criteria for establishing the operations timetable shown in Table 16.2.7_1 were adopted in

order to select the necessary equipment.

Table 16.2.7_1

Mara Rosa Project

Operations Timetable Criteria

Items Unit Value

Number of scheduled hours per day h 24

- Meal breaks h 1.5

- Shift changes h 1.0

Number of scheduled hours per day h 21.5

Scheduled days per year day 365

Idle days per year due to severeweather conditions day 5

Working days per year day 360

Number of scheduled hours per year h 7,740

Scheduled maintenance (hours per year)

- Lubricant changes h 155

- Preventive maintenance h 330

- Overall repairs h 150

Scheduled Sub-Total h 635

Non-scheduled h 155

Total Maintenance h 790

Estimated mechanical availability % 90%

Operational delays

- Blasting h 100

- Fuelling h 100

- Overall operations h 50

- Unspecified h 100

Total Operational delays 350



Available hours per year h 6,600

Use of availability % 80%

Actual hours worked per year h 5300

Actual operational hours per day h 14.72

Effective operational use/deployment % 68%

The total hours that are effectively available were calculated with respect to 360 operational

days per year (considering 5 days idle time due to severe weather conditions) and 21.5 hours

per day (7,154 hours per year) distributed over three shifts of eight hours. There are a total of

5,300 effective operational hours per year.

16.2.8 Drilling Equipment and Productivity

A generic grid was prepared for drilling and blasting of the ore and fresh waste rock, taking

into account the competence of the rock and use of a 4" drill bit for the ore and a 5" drill bit for

the waste rock. The saprolite located in the upper section of the deposit does not require

drilling and fragmentation.

Mining operations carried out on 10 m benches was considered standard procedure for the

project. However, with respect to the mining operations being performed on 5 m benches, a

certain amount of flexibility was granted with respect to equipment requirement estimates and

selective mining operational costs at the contact areas of the mineralized zones. This should

affect 30% of the total tonnage of ore in the pit.

The main parameters that were considered are depicted in Table 16.2.8_1.

Table 16.2.8_1

Mara Rosa Project

Drilling and Blasting Parameters

Ore Waste Rock

Hole diameter (inches) 4.00 5.00

Explosive diameter (inches) 4.00 4.00

Rock density (g/cm3) 2.73 2.73

Explosive Density (g/cm3) 0.94 0.94

Bench Height (m) 5.00/10.00 10.00

Burden (distance between rows) (m) 3.00 3.50

Spacing (distance between holes) (m) 6.00 6.50

Sub-drilling (m) 0.40/0.70 0.70

Length of hole (m) 5.40/10.70 10.70

Collar (m) 0.70/1.40 1.40

Explosive column (m) 4.70/9.30 9.30

Specific Charge (kg/m) 6.75 6.75

Explosive Charge per Hole (kg) 31.35/62.70 62.70

Hole Volume (m3) 90.00/180.00 228.00

Total Blasted Rock (t) – 5 m bench 11,057.00 27,948.38

Total Blasted Rock (t) – 10 m bench 22,113.00 55,896.76

Powder Ratio (g/m3) 348.32 269.89



The performance level of the drilling equipment was based on the experience of suppliers.

The design parameters and operational levels mentioned above were not used for this project

because they were considered optimistic for these operations.

As a result of these productivity calculations, the annual capacity was estimated for each type

of drill and for each period of mining operations as depicted in Tables 16.2.8_2 and 16.2.8_3.

The tables also present the production scenarios for each of the seven years during which the

mine will be active. The foreseen production capacity is constant during this time period.

A productivity of 30 m/h was applied for the 5 inch drill and 15 m/h for 4 inch drill (this last was

underestimated because of reverse circulation and larger displacements in the case of 5 m

benches). These parameters were estimated as a function of the competence level of the

rock.

Table 16.2.8_2

Mara Rosa Project

Drilling Performance in Ore

Ore Drill Unit Year

1 2 3 4 5 6 7

Diameter mm 101.6 101.6 101.6 101.6 101.6 101.6 101.6

Volume per meter drilled m3/m 78.5 78.5 78.5 78.5 78.5 78.5 78.5

Tons per meter drilled t/m 214 214 214 214 214 214 214

Drilling efficiency % 42 42 42 42 42 42 42

Penetration rate m/h 15 15 15 15 15 15 15

Tons per hour t/h 689 689 689 689 689 689 689

Mechanical Availability % 87 87 87 87 87 87 87

Factor Utilized % 48 48 48 48 48 48 48

Hours per shift h 8 8 8 8 8 8 8

Shifts per day ea 3 3 3 3 3 3 3

Days per year d 360 360 360 360 360 360 360

Annual production kt 2,298 2,404 2,451 2,440 2,517 2,444 2,361

Table 16.2.8_3

Mara Rosa Project Drilling Performance in Waste

Waste rock drill Unit Year

1 2 3 4 5 6 7

Diameter mm 127 127 127 127 127 127 127

Volume per meter drilled m3/m 56.9 56.9 56.9 56.9 56.9 56.9 56.9

Tonnes per meter drilled t/m 163 163 163 163 163 163 163

Drilling efficiency % 42 42 42 42 42 42 42

Penetration rate m/h 15 15 15 15 15 15 15

Tonnes per hour t/h 1,915 1,915 1,915 1,915 1,915 1,915 1,915


Factor Utilized % 48 48 48 48 48 48 48




Shifts per day ea 3 3 3 3 3 3 3

Days per year d 360 360 360 360 360 360 360


This rate of productivity is then utilized to estimate the quantity of equipment necessary per

year at the Posse Mine. With this in mind, the drilling work was distributed among two types of

drill rigs, as shown in the Table 16.2.8_4.

The drilling operations at the Posse Mine will be conducted entirely with diesel-powered drill

rigs, which are fully capable of providing the necessary rate of production throughout the life

of the mine.

At the Posse Mine, the four-inch drills will be deployed at the ore benches, while the five-inch

drills will operate on the waste rock.

The five-inch diameter drills will conduct approximately 70% of the operations; the four-inch

drills will perform the remainder of the work. Considering this distribution of work and the

parameters listed above, the number of drill rigs that would be necessary each year was

estimated. Table 16.2.8_4 summarizes these figures.

Table 16.2.8_4

Mara Rosa Project Required Number of Drill Rigs

Equipment Unit Year

1 2 3 4 5 6 7

Ore Drill ea 2 2 2 2 2 2 2

Waste Rock Drill ea 3 4 4 4 4 2 2

These numbers were weighted for each operational index; as a result, they represent the total

number of machines necessary to carry out the mining plan.

16.2.9 Loading Equipment and Productivity

The productivity rates of the loading equipment were estimated based on the above-

mentioned operational parameters and detailed estimates of how much time would be

required for loading the material.

These calculations were performed for unaltered rock, having an average sponge density of

1.95 t/m3 and 3% moisture content, and for altered rock (saprolite), having a sponge density

of 1.71 t/m3 and a moisture content of 3% as well.

It was assumed that all of the saprolitic material would be excavated, loaded, transported and

piled up during the pre-stripping phase. Tables 16.2.9_1 and 16.2.9_2 display the productivity

calculations for each unit operating on waste, and ore and altered rock respectively.



Table 16.2.9_1

Mara Rosa Project Productivity Data for Waste Rock

HYDRAULIC EXCAVATOR EX 1200 Waste Rock

Unit Year

1 2 3 4 5 6 7

Bucket capacity yd3 13.73 13.73 13.73 13.73 13.73 13.73 13.73

Bucket capacity m3 10.50 10.50 10.50 10.50 10.50 10.50 10.50

Fill factor % 85 85 85 85 85 85 85

Freely settled density t/m3 1.95 1.95 1.95 1.95 1.95 1.95 1.95

Moisture content % 3 3 3 3 3 3 3

Dry tons per bucket t 17.4 17.4 17.4 17.4 17.4 17.4 17.4

Tons per bucket t 17.9 17.9 17.9 17.9 17.9 17.9 17.9

Cycle time min 1.20 1.20 1.20 1.20 1.20 1.20 1.20

Passes per truck p 6 6 6 6 6 6 6

Dry tons per truck t 92.20 92.20 92.20 92.20 92.20 92.20 92.20

Loading time min 2.27 2.27 2.27 2.27 2.27 2.27 2.27

Waiting and spotting time min 1.33 1.33 1.33 1.33 1.33 1.33 1.33

Total loading time min 3.61 3.61 3.61 3.61 3.61 3.61 3.61

Effective loading @ 100 (th) t/h 2,256 2,256 2,256 2,256 2,256 2,256 2,256

Loading efficiency % 68 68 68 68 68 68 68

Tons per operating hour t/h 2,256 2,256 2,256 2,256 2,256 2,256 2,256


Utilization factor % 80 80 80 80 80 80 80


Shifts per day t 3 3 3 3 3 3 3

Days per year d 360 360 360 360 360 360 360




Table 16.2.9_2

Mara Rosa Project Productivity Data for Ore and Altered Rock

HYDRAULIC EXCAVATOR EX 1900

Unit Year

1 2 3 4 5 6 7

Bucket capacity yd3 7.58 7.58 7.58 7.58 7.58 7.58 7.58

Bucket capacity m3 5.8 5.8 5.8 5.8 5.8 5.8 5.8

Fill factor % 85 85 85 85 85 85 85

Sponge density t/m3 1.95 1.95 1.95 1.95 1.95 1.95 1.95

Moisture content % 3 3 3 3 3 3 3

Dry tons per bucket t 9.6 9.6 9.6 9.6 9.6 9.6 9.6

Tons per bucket t 9.9 9.9 9.9 9.9 9.9 9.9 9.9

Cycle time min 0.80 0.80 0.80 0.80 0.80 0.80 0.80

Passes per truck p 5 5 5 5 5 5 5

Dry tons per truck t 43.7 43.7 43.7 43.7 43.7 43.7 43.7

Loading time min 1.95 1.95 1.95 1.95 1.95 1.95 1.95

Waiting and spotting time min 1.33 1.33 1.33 1.33 1.33 1.33 1.33

Total loading time min 3.28 3.28 3.28 3.28 3.28 3.28 3.28

Loading efficiency % 68 68 68 68 68 68 68

Tons per operating hour t/h 1,246 1,246 1,246 1,246 1,246 1,246 1,246


Utilization factor % 80 80 80 80 80 80 80

Shifts per day t 3 3 3 3 3 3 3

Days per year d 360 360 360 360 360 360 360


A 10% loss in efficiency was applied to the hydraulic excavator at the Posse Mine for

selective mining of the ore at the contact areas of the mineralized zones. A similar parameter

was applied to the drill rigs.

The number of loading equipment that would be necessary was estimated with respect to the

required recovery rate of the stockpile. In this case, to arrive at an estimate, it was assumed

that front loaders, which have a slightly higher loading efficiency, would do all of the loading.

As such, an estimate of 1,607t/operating hour was achieved (approximately 10% better than

the loading that occurs at the mine).

Hydraulic excavators will be utilized at the Posse Mine to excavate ore and rock waste until

both materials have been completely removed.

The hydraulic excavators will load all of the rock waste; the front loaders may be used to load

the ore on a contingency basis.



Taking into account the above-mentioned work distribution and parameters, an estimate of

the amount of vehicles required per year was made.

Transport Distance Calculations

Transport distances were measured in the profiles of the pit and waste dumps, for each

destination and for each year of the mine. A summary of the distances that were obtained for

each of these cases can be viewed in Table 16.2.9_3.

Table 16.2.9_3

Mara Rosa Project

Transport Distances (m)

Material Year

0 1 2 3 4 5 6 7

Waste Rock pile (m) 883 1,094 1,314 1,564 2,667 2,745 3,117 3,131

Ore pile (m) 672 921 1,021 1,399 1,775 2,185 2,627 2,994

These figures do not take into account the recovery of the low-grade ore from the stockpile.

Truck Velocity

Table 16.2.9_4 shows the average velocities used to estimate the travel time for each pit,

period and type of material.

Table 16.2.9_4

Mara Rosa Project

Truck Velocities

In the pit (km/h) Outside of pit (km/h)

Hauling uphill loaded @ 10% 11 11

Hauling downhill empty @ 10% 35 35

Hauling downhill loaded @ 10% 30 30

Hauling uphill empty @ 10% 25 25

Hauling Empty on horizontal terrain 30 40

Hauling Full on horizontal terrain 30 40

These speeds were applied to every transport profile that was obtained in order to calculate

the total travel time. In order to obtain the values that were adopted, a Coffey Mining

simulation program was used to perform cycle simulations. The total number of simulated

truck hours for each period was obtained with the methodology applied by Coffey Mining.



Detailed information with respect to the number of cycles performed per year and by type is

presented in Fonseca and Horta (2011).

Truck Productivity

Truck productivity was calculated for the entirety of the loading and production period for each

of the pits and for the saprolite. It depended on the travel time and other factors within the

cycle which time durations are constant (loading by type of loading equipment used, dump,

positioning, etc.).

Details regarding calculations of truck productivity as a function of excavator type and material

type are presented in the section: Mine Schedule.

A summary of the main parameters used in calculating truck performance is contained in

Table 16.2.9_5.

Table 16.2.9_5

Mara Rosa Project

Parameters Used in Calculating Truck Productivity

Truck - Excavator Unit Ore Waste

Nominal Truck Capacity t 45 95

Moisture content % 3 3

Number of Passes w/ Hydraulic Excavator - 5 6

Effective Load w/ Hydraulic Excavator t 9.6 17.4

Loading Time w/ Hydraulic Excavator min 1.9 2.3

Unloading and Manoeuvring min 1.6 1.3

Wait Time min 0.5 0.5

Truck Efficiency % 68 68

Truck Hours Necessary

The number of trucks that are necessary was estimated by dividing the total amount of

material that needs to be transported during the time required based on the capacity of the

trucks for each combination of pit, time period and loading equipment (Table 6.17).

The quantity of truck hours was calculated with respect to year, type of material and loading

equipment, dividing the tonnage that needs to be transported by the hourly productivity of

each combination.

A summary of the truck hours, truck productivity and how many trucks are necessary is

presented in Tables 16.2.9_6 and 16.2.9_7.



Table 16.2.9_6

Mara Rosa Project

Transport Time and Quantity of Equipment Required for Ore truck

Equipment Unit Year

1 2 3 4 5 6 7

Tonnage kt 7,454 8,203 9,918 11,415 13,505 14,924

Total Cycle Time min 7.0 8.1 8.5 10.2 11.8 13.6 15.6

Hourly Productivity t/h 368 334 317 265 228 198 173

Total Productivity Hours

h 534 7,088 7,821 9,528 11,027 13,105 14,536

N° of Trucks ea 1 2 2 2 3 3 3

Table 16.2.9_7

Mara Rosa Project

Transport Time and Quantity of Equipment Required for Waste truck

Equipment Unit Year

0 1 2 3 4 5 6 7

Tonnage kt 11,333 18,684 25,054 25,316 24,369 23,650 10,083 2,815

Total Cycle Time min 8.0 8.9 9.9 11.0 15.8 16.2 17.8 17.9

Hourly Productivity t/h 713 639 576 519 360 352 320 319

Total Productivity Hours

h 15,892 29,243 43,468 48,806 67,726 67,157 31,529 8,830

N° of Trucks ea 4 6 9 10 13 13 6 2

Taking into account the above-mentioned work distribution and parameters, an estimate of

the number of vehicles required per year was made. The maximum number of trucks

required is 16 (13 trucks for transporting waste rock and 3 for transporting ore) for year 4.

During the last two years of operation (6 and 7), fewer trucks will be necessary because of the

reduced tonnage of waste rock that will need to be transported.

16.2.10 Ancillary and Support Equipment

Ancillary equipment was selected by considering the size and type of the main loading and

transport fleet, the geometry and size of the pits and the number of roads and waste rock

piles that will be operating at the same time.

The number of ancillary and support equipment units is summarized in Table 16.2.10_1.



Table 16.2.10_1

Mara Rosa Project

Number of Ancillary and Support Equipment Required

Equipment Unit Year

1 2 3 4 5 6 7

Bulldozer ea 2 2 2 2 2 2 2

Moto Grader ea 1 1 1 1 1 1 1

Water Truck ea 1 1 1 1 1 1 1

Truck Loading Crane ea 1 1 1 1 1 1 1

Fuel Truck ea 1 1 1 1 1 1 1

Lube Truck ea 1 1 1 1 1 1 1

Crane - 50t capacity ea 1 1 1 1 1 1 1

16.2.11 Total Fleet Required for the Mine

The total quantity of equipment necessary according to each year of production is

summarized in Tables 16.2.11_1 and 16.2.11_2.

Table 16.2.11_1

Mara Rosa Project

Main and Auxiliary Equipment Necessary

Equipment Unit Year1 2 3 4 5 6 7

Main Equipment

Drilling - Ore ea 1 2 2 2 2 2 2

Drilling - Waste ea 2 3 4 4 4 4 2

Front loader ea 2 2 2 2 2 2 2

Hydraulic Excavator - Ore ea 1 2 2 2 2 2 2

Hydraulic Excavator - Waste ea 1 2 2 2 2 2 2

Trucks - Waste, ea 3 6 9 10 13 13 6

Trucks - Ore ea 1 2 2 2 3 3 3

Auxiliary Equipment

Bulldozer ea 2 2 2 2 2 2 2

Moto Grader ea 1 1 1 1 1 1 1

Water Truck ea 1 1 1 1 1 1 1



Table 16.2.11_2

Mara Rosa Project

Necessary Support Equipment

Equipment Unit Year

1 2 3 4 5 6 7

Truck Loading Crane ea 1 1 1 1 1 1 1

Service Truck ea 2 2 2 2 2 2 2

Flat-bed Truck ea 1 1 1 1 1 1 1

Crane - 50t capacity ea 1 1 1 1 1 1 1

16.2.12 Establishment of Equipment Lifetime

The lifetime of each piece of mining equipment was established in total hours of operation

according to data from other operations that use similar equipment and from information that

was supplied by the equipment manufacturers.

16.3 Description of Pit Operation and Infrastructure

The support facilities, which are unique to the mine, include the drainage system and the

electrical supply grid that is necessary to provide power to the pit’s drainage system. Other

support facilities include those necessary for explosives storage, for the housing of

computerized control systems and those necessary for mine operations planning.

16.3.1 Pit Drainage

Water enters the pit of the Posse Mine directly through precipitation, superficial run-off and

through groundwater. This volume of water must be pumped out of the pit to the water

reservoir. The majority of this water enters the pit during the wet season, which begins in

October and continues through April.

During the dry season, only groundwater needs to be removed from the pits, the greater part

of which collects automatically at the lowest section of these structures.

It rains during the majority of days during the rainy season, making it necessary to pump this

water out of the pits in order to keep the mining fronts operational. To minimize the flow of

water into the pit at the Posse Mine, a system of channels and containment basins will be built

to divert and contain the flow of water on both sides (northwest and southeast) of the pit. The

most important containment basins are located on the southeast side of the pit. The system of

containment basins and channels on the southeast side of the Posse pit will collect up to two-

thirds of the water that would normally flow into the pit during the wet season. Provided that

this water does not come into contact with the material that is inside the mine, it may be

diverted directly into the Araras Stream.



The system of pumps was designed to keep the pit free of water at precipitation levels of up to

300 mm (highlighted in the original) per month. Although the mining fronts will still be

operational during the wet season, mining operations will be less intense at this time. In the

event of heavy rains, the ore that is fed into the plant will be supplemented with ore from the

stockpile. The total amount of pit drainage that is foreseen as of year 3 is 600 m3/h, which will

require 4,105 kWh of electricity. Plans specify that this water be sent to the water reservoir to

be used in ore processing.

16.3.2 Providing Electricity to Pit Operations

A transmission line will be installed adjacent to the final pit boundary that will be responsible

for providing power to the pump system. The supply of electricity to the drainage pumps will

be necessary at all times; however, only one transmission line will be available on the south

eastern side of the pit to fulfil this requirement.

16.3.3 Storage and Preparation of Explosives

The extraction of the blasted rock will be done by Amarillo’s staff and supervised by the

blasting contractor. The supply of explosives will include a combination of ANFO (Ammonium

Nitrate/Fuel Oil) and emulsion cartridges. The smaller area will be used for storing twine,

dynamite and other accessories.

This area will be located far enough from the mine and support facilities so as to be in

compliance with Brazilian safety regulations.



17 RECOVERY METHODS

17.1 Basic and General Criteria

17.1.1 Bases and Units Used

The full Plant Design study in Portuguese was completed by Onix (2011). The following

section is an abridged version of that report for inclusion in this NI43-101 report.

As noted in Section 2.6, all units are metric SI units. Costs have been derived from various

sources but all are presented in US$ after conversion at the stated exchange rates. For Brazil

Reais an exchange rate of R$1.9 = US$1.

17.1.2 Definition of Capacity

The plant has been designed to treat a nominal annual tonnage of 2.5 Mtpa. In the case of the

crushing circuit an overall utilization rate of 75% has been used which means that the crusher

circuit throughput is 381 tph. A value of 90% overall utilization has been used for the milling

circuit which implies an hourly throughput of 317 t.

17.1.3 Project Base

Table 17.1.3_1

Mara Rosa Project

Data Sources for Plant Design

Assumed A

Advice from Coffey B

Advice from Amarillogold C

Calculation D

Supplier Advice E

Estimated F

Engineer Experience G

Statutory Authority H

Testwork T

To be advised TBA

Not Applicable NA

17.1.4 Geography

Table 17.1.4_1

Mara Rosa Project

Physiographic Dat

Location 7 km. from the Mara Rosa town

Location of town 11 km. west of BR-153 in the northern part of Goiás

Altitude Approximately 420 m above sea level

Vegetation type Previously tree covered, (Cerrado), but now mostly pasture



Table 17.1.4_2

Mara Rosa Project

Climate Data

Mean Daily minimum temperature 22 degrees centigrade H

Mean Daily maximum temperature 32 degrees centigrade H

Yearly Average Rainfall 1679 mms. H

Wet days per year 114 days H

Mean 3 pm Relative Humidity 66% H

Mean 3 pm wind speed 1 m/sec. H

17.1.5 Metallurgical Testwork

During plant operations from 1992 to 1995, Western Mining Corporation were aware that

recoveries were declining as the pit was deepening and less oxide ore and more sulphide ore

(with tellurides) was being processed. The recovery in the final month of production before

closure due to the low gold price was only 83%.

A large number of testwork campaigns have been carried out over the years, mostly

misguided as the proponents did not have the necessary experience to design programmes to

allow for the presence of the tellurides. High extractions were only achieved (Western Mining

Laboratories and Testwork Technologies) when oxidation was applied in the form of high

additions of calcium hypochlorite, implying a very expensive process route.

Under the guidance of metallurgical consultants with personal experience of deposits

containing tellurides (in particular, the Finiston mine in Kalgoorlie operated by Kalgoorlie

Consolidated Gold Mines), a programme involving oxidation on a more practical level in terms

of costs was formulated (see Section 13 for details).

As a result the laboratory testwork indicated that recoveries of 93% could be obtained from

the two main ore types (Main and Hanging Wall). It is considered that this applies to more

than 97% of the gold content of the deposit. The ore type FW, of lower grade, with less than

3% of proven or probable reserves gave recoveries in the order of 86% under the same

conditions.

The final process route is again applying strong oxidation but by milling to a P80 of 45 μm and

exposing the ground mineral to oxidation of the pulp with the injection of low grade oxygen

gas (delivered by a cheap PSA plant) for 12 hours at a pH of 12, an economic process route

with a high recovery in the subsequent cyanidation stage has been achieved.

The ball mill work index is at 13 kWh/t, considered reasonable and therefore the fine grinding

does not involve excessive costs. The PSA oxygen plant will also consume some additional

energy (325 kW) and a calculation of the overall extra energy cost per tonne of ore results in a

value of 7 kWh/t processed. This includes the extra total agitation time of 36 hours against the

more usual 24 hours.



The final design parameters derived from the testwork are presented in Table 13.10.1.

17.2 Process Description

17.2.1 Overview of Process

The plant consists of a conventional crushing circuit (including tertiary crushing) followed by

primary and then secondary milling in closed circuit with cyclones. The final pulp at a P80 of

45 µm is pre-oxidized in agitated tanks using oxygen gas from a PSA oxygen plant at a high

pH of 12 for a total of 12 hours to oxidize tellurides to enable successful cyanidation of the

gold. The pulp is contacted with cyanide and activated carbon in a typical CIL circuit of six

agitated tanks for a total of 24 hours.

Loaded carbon is extracted daily from the CIL circuit and processed in a typical Zadra style

elution circuit at up to 140° C with a 4 tonne capacity. The eluted solution is passed

continuously the electrowinning cells until efficient desorption has been achieved. At intervals

the gold is removed from the cells and smelted into Doré bars for sale. The activated carbon

is regenerated in a gas fired rotating kiln before being sent back to the CIL circuit.

The tails form the CIL circuit is thickened to recover some of the solution before the thickened

pulp is subjected to detoxification with SO2/air and a copper sulphate catalyst to destroy free

cyanide before being pumped to a tailings storage facility. The supernatant from TSF is

recirculated to the plant under conditions of zero discharge.

An overview of the material characteristics is presented in Table 17.2.1_1.



Table 17.2.1_1

Mara Rosa Project

Material Characteristics

Units Number Reference

Feed Grade, Au, in situ g/t 1.77 B

Mine Dilution % 3 B

Final ROM grade g/t 1.72 D

Design Moisture Content % 3 B

Gold Recovery

Gravity % 15 A

CIL %, (of feed) 77 T

Total % 92 T

Annual Throughput t 2,500,000 B

Total Average Gold Recovery grams/day 10,838 D

Total Average Gold Recovery troy ounces/annum 127,188 D

Unconfined Compression Strength

Range Mpa 59-211 A

Design Mpa 140 A

Crushing Work Index

Minimum kWh/t 4.5 A

Maximum kWh/t 10.5 A

Average kWh/t 6.5 A

Design kWh/t 10.5 A

Rod work Index kWh/t 13.4 T

Ball work Index kWh/t 13.0 T

BWI Design kWh/t 13.5 G

Abrasion Index kWh/t 0.343 T

Ore Specific Gravity t/m3 2.73 T

Settling Characteristics t/m2/h 0.5 A

Water Specific Gravity t/m3 1.0 A

Ore Bulk Density t/m3 1.95 A

17.2.2 Crusher Circuit

Ore from the open pit is dumped through a 800 mm x 800 mm grizzly into the jaw crusher

feed hopper (ROM bin) with capacity of 140 t. The grizzly will be sloped in such a way so that

a wheel loader can easily withdraw oversize to be dealt with using a portable hydraulic rock

breaker. There is an opportunity to stockpile ROM ore near the ROM bin for later treatment

by feeding with a wheel loader. The grizzly is designed in two halves and fits onto the top of

the bin in such a way as it can be easily lifted off when blockages between the grizzly and the

vibrating feeder occasionally occur.

A vibrating feeder with attached grizzly (with capacity in excess of 500 tph) with variable

speed motor of approximately 22 kW draws the ore out from the ROM bin and discharges

oversize into a jaw crusher 1400 x 830 mm (Telsmith HD 1400 or equivalent) which crushes

to a P80 of less than 140 mm.



Fine atomizing sprays keep down the level of dust.

The scalped fines and the product from the crusher fall onto the same 1000 mm crusher

discharge belt running at 1.0 m/s which feeds a second similar belt running at 1.2 m/s which

feeds a stockpile. Ore is withdrawn from underneath the stockpile by one of two variable

vibrating feeders and transferred by conveyor belt to a double decked vibrating screen. The

undersize at a size of 100% passing 19 mm discharges onto a conveyor belt (which also

receives product from both cone crushers). The oversize (from both the screens 19 mm. and

75 mm.) is discharged into a secondary cone crusher from whence the product joins the

undersize from the primary screen as well as the product from the tertiary cone crusher. This

total product is then transferred to a further belt which feeds a further double deck secondary

screen (19 mm. and 38 mm.) Undersize from this screen is the final -19 mm product which is

transferred by conveyor to the mill feed silo. Oversize from this screen is conveyed to a

further stockpile. From under this stockpile three small variable speed vibrating chutes feed a

conveyor which feeds the tertiary cone crusher to complete the circuit..

A weightometer is located on the final 1000 mm conveyor belt feeding the fine ore bin.

An overview of the crushing criteria is presented in Table 17.2.2_1.

Table 17.1.3_1

Mara Rosa Project

Plant Crushing Criteria


Operating Schedule

Annual Throughput tpa 2,500,000 C

Crushing Circuit stages 3 B

Operating Days per annum days 365 D

Total Operating Hours hours 8760 D

Effective Utilization % 75 G

Effective Operating Hours hours 6570 D

Required Crushing Rate tph 381 D

ROM Bin

Live Capacity t 140 G

Capacity at required crushing rate min 22.1 D

Static grizzly Bar Spacing mms 800 G

Primary Feeder

Type Vibrating feeder/grizzly G

Size m 1.07 x 4.27 E

Maximum Capacity tph 500 E

Primary Crusher

Type Single Toggle Jaw Crusher Size

Telsmith HD 1400 or equivalent E

Maximum feed size mms 800 E

Product Size, (P100) mms 185 A

Product Size, (P80) mms 140 A

Design Closed setting mms 125 A




Conveyors, (crushing circuit)

Number 11 E

Size, (width) mms 1000 E

Length, various m 8-34 E

Vibrating Feeders

Number in circuit 5 E

Primary Vibrating Screen

Type Double deck 75mm and 19mm E

Size m 1.8 x 4.9 E

Secondary Cone Crusher

Type Telsmith 52SBS or equivalent E

Motor kW 300 E

Design CSS mms 50 G

Secondary Vibrating Screen

Type Double deck 38mm and 19mm E

Size m 2.4 x 6.1 E

Tertiary Cone Crusher

Type Telsmith 52SBS or equivalent E

Motor kW 300 E

Design CSS mms 13 G

Fine Ore Storage Bin

Type Cylindrical, steel with overflow G

Live Capacity t 3,500 G

Equivalent milling Capacity h 11 D

Reclaim Method Feed conveyors G

Number 4 G

Reclaim Feeder Size mms 600 G

Reclaim Feeder Length m 25 A

Reclaim Feeder Maximum Capacity tph 400 G

Ball Mill Feed Conveyor

Size, width mms 1000 G

Size, length m 83 A

Lime Addition System

Vertical conical silo with feeding system, additions using adjustable screw conveyor

17.2.3 Mill Feed Circuit

The fine ore bin will have a total live capacity of 3,500 t and when full automatically overflows

to open stockpiles. This material can be recycled to the final crushing belt using wheel loaders

so as to avoid mill down-time during prolonged crusher interruptions. The fine ore (100% less

than 19 mm with a P80 of 12 mm) is reclaimed from the fine ore bin using four 0.6 m x 25 m

belt feeders fitted with variable speed drives at a total rate of 317 dry tons per hour (nominal

feed rate) and is discharged onto a 1,000 mm wide ball mill feed conveyor belt.

Lime is added to this same belt using a variable speed screw feeder from a silo mounted to

the side of the belt. Consumption of lime is anticipated to be 15 tpd and will arrive daily in a

tanker which will transfer the lime pneumatically. Additions of lime will be made in order that

the pH in the pre-oxidation tanks is maintained at 12.0. The lime silo will have a capacity of



60 t and will be equipped with filtering devices so that lime will not escape to the atmosphere.

The area around the lime silo will be concreted and have a sump pump to return spillage to

circuit.

17.2.4 Mill Circuit

The mill feed conveyor will feed ore into the top of a detachable spout trolley lined with wear

plates which in turn feeds the primary fixed speed ball mill with basic approximate size of

5.5 m x 8 m with a power rating of 4 MW. The mill feed spout trolleys for each mill sit on rails

on the mill working platform and can be pushed back out of the way if necessary. The mill

platform is large enough for a liner handler to be installed when re-lining is called for, as well

as other ancillary equipments and a control room. Water is added tangentially into the spout

to help wash material into the mill.

The circuit has two ball mills with similar dimensions with the difference that the first mill (the

primary mill) will have larger mill balls and that this mill will be in open circuit. The feed to the

secondary ball mill will consist of cyclone underflow and will use smaller mill balls for a finer

grind.

Balls (70 mm. for the primary and 25 mm. for the secondary mill) will be loaded via a two

tonne kibble so that they can be fed into the feed chutes of the mills using an electric hoist

which lifts the ball kibble and is conveniently sited with a swivel arm to carry out the task. This

hoist will also be used to carry out maintenance on the cyclones. The mills are mounted on

concrete foundations so that there is a 2.5 m head room from the bottom of the mills to the

concrete floor. Sufficient access will be provided in the design so that ore spills at the feed

end of the mill can be cleared with a Bob-Cat or equivalent. Also the whole mill area (including

pumps) is bunded so that any spillage is confined. Sump pumps are installed so that spillage

can be returned to the circuit.

The pulp exits the each of the ball mills by way of a polyurethane trommel (with sprays) which

ensures that used balls and oversize are separated out and fall into a bunker whilst the fines

(as a pulp) fall into the mill distribution chute. This distribution chute feeds either of two pump

boxes with appropriate dart valves which are connected to two horizontal 12 x 10 centrifugal

cyclone feed pumps (one stand-by). The pump boxes are rubber lined (as are all similar

boxes around the plant). The cyclone feed pumps pump the pulp from the hopper to a cyclone

nest of 8 units of 380 mm cyclones, (five5 on line and three spare, three with automatically

remote controlled feed valves). Like all the pulp pump lines, no part is horizontal, maintaining

a 2% minimum angle. Above the cyclone feed pumps exists a monorail to facilitate

maintenance, (as is the case of all the large horizontal pumps on the plant). The underflow

from these cyclones is directed into the secondary ball mill feed using launders whilst the

overflow (P80 -45 µm) goes to the pre-oxidation/leach train with a pulp density of 45% w/w

solids.

An overview of the grinding and classification circuit is presented in Table 17.2.4_1.



Table 17.2.4_1

Mara Rosa Project

Grinding and Classification Parameters


General

Type Milling with ball mills and classification with cyclones

B

Design Bond Index kWh/t 13.5 T

Abrasion Index, design 0.343 T

Ore Specific Gravity t/m3 2.73 T

Water Specific Gravity t/m3 1.0 A

Feed size, (P80) mms 12.0 G

Product Size, (P80) microns 45 G

Operating Schedule

Annual Nominal Throughput t 2,500,000 B

Days per Year days 365 D

Hours per day hrs 24 D

Overall Utilization % 90 G

Effective Operating hours hrs 7884 D

Required Treatment Rate tph 317 D

Average daily production tpd 7610 D

Primary Mill

Type Overflow, Open Circuit Size

(inside liner diameter x EGL) m 5.5 x 8 E

Power at Pinion kW 3,800 E

Motor draw Power kW 4,000 E

Mill Speed fixed % Cs 75 E

Liner Material Rubber G

Design Ball Charge % 33 E

Make Up Ball Size mms 70 E

Mill Discharge Density % 75 G

Discharge Screen type trammel G

Discharge Screen Aperture mms Two parts 6 x 20 and 14 x 20

G

Screen Discharge Material Polyurethane G

Secondary Mill

Type Overflow, Closed Circuit with cyclones

G

Size (inside liner diameter x EGL) m 5.5 x 8 E

Power at Pinion kW 3,800 E

Motor draw Power kW 4,000 E

Mill Speed fixed % Cs 75 G

Liner Material Rubber G

Design Ball Charge % 36 E

Make Up Ball Size mms 25 E

Mill Discharge Density % 72 G

Discharge Screen type trammel G

Discharge Screen Aperture mms Two parts 6 x 20 and 14 x 20

Screen Discharge Material

Polyurethane




Cyclone Feed Pump

Type Rubber lined horizontal centrifugal

G

Size Warman 12 x 10 or equivalent E

Flow m3/hr 931 G

Material pulp, maximum diameter 14 mms.

G

Density S.G of solids 2.73, pulp density 62% w/w

G

Pressure kpa 300 F

Classification

Cyclones Krebs Gmax 15 or similar E

Size mms 380 E

Overflow density %w/w 45 C

Cyclone Feed Flow m3/hr 931 G

Cyclone Feed Density %w/w 62 G

Underflow density %w/w 76 G

Cyclone Recirculating Load % 200 G

Operating Pressure kpa 150 F

Product Size, (P80) microns 45 B

Sump Pump

Type Vertical Pump Warman 65QVSP or similar

G

17.2.5 Gravity Circuit

As previously mentioned the ball mills have an outlet trommel and this has slots of 14 mm x

100 mm. However the initial section has apertures of only 6 mm. Under this portion of the

trommels a separate chute collects the fines and diverts this to a 6 x 4 steel lined heavy duty

horizontal centrifugal gravity feed pump situated on the lower mill bunded floor. There exist

dart valves which can isolate this pump. Any overflow from this initial chute section (either

because of reduced pumping capacity of the gravity feed pump or it is off line) overflows this

initial chute section into the main mill discharge distribution chute.

The gravity feed pump pumps up to a vibrating screen rated at 200 tph with 2.4 mm

polyurethane slotted screen. The oversize joins the launder containing cyclone underflow

which feeds back into the Ball mill. The undersize feeds a KC-XD40-MS Knelson concentrator

or a Falcon SB2500. This is an automatic device and when it goes off-line (every 2 hours)

then its feed gets diverted back to the same launder feeding back to the Ball mill. The same

electric hoist used for the balls will also be used for maintenance to the vibrating screen and

the concentrator.

The concentrate discharging from the Knelson (at two hourly intervals) is fed into a holding

hopper for the intensive cyanide leach system (Acacia or Gekko). This system is a semi-

automatic leaching system which will leach any free oxidized gold in an intensive leach (high

cyanide and under highly oxidizing conditions). The leaching occurs in batch form,

(approximately 500 kg) with one batch being treated every 24 hours. The leach liquor is



transferred to the gold room where it is electrolysed in its own cell (approximately 600 mm x

600 mm). This cell is cleaned out weekly. The leached residue is pumped back to the launder

which receives cyclone underflow.

An overview of the gravity circuit is presented in Table 17.2.5_1.

Table 17.2.5_1

Mara Rosa Project

Gravity Circuit Parameters


Gravity Feed Pump

Type Warman 6 x4 or similar E

Pulp maximum particle 6 mms,

201 m3/hr G

Gravity Separation Screen Sizetech or similar G

Feed Ball Mill Discharge

Design Feed Solids tph 200 C

Diluted Feed Density % w/w 61 G

Design Screen Flow m3/h 201 C

Type Horizontal Vibrating with

polyurethane panels G

Nominal Screen Size m 1.5 x 3.6 G

Screen Aperture mms 2.4 G

Screen Oversize Destiny Ball Mill Feed G

Screen Undersize Destiny Centrifugal Concentrator G

Centrifugal Concentrator

Feed Source Gravity Separation Screen

Undersize G

Design Feed Solids tph 200 C

Diluted Feed Density % w/w 61 G

Total Pulp Flow m3/h 201 C

Type of Concentrator Knelson XD40 or Falcon

SB2500 C

Total Fluidizing water m3/hr 35 E

Concentrate Recovery Cycle Time hr 2 C

Concentrate recovery each cycle kg 40 E

Total Cycles per day 12 C

Total Concentrate Production kg/day 480 D

Concentrate Destination Intensive Cyanidation

Tailings Destination Ball Mill Feed

Intensive Cyanidation

Type Gekko or Acacia C

Concentrate Treatment Rate kg/day 480 D

Design Global Recovery Rate % 15 A

Design Gravity Gold grams/day 1962 D

Design Recovery grams/hr 82 D

Concentrate Gold Grade g/t 4089 D

Solution Tank Volume m3 1.5 E

Final Liquor solution grade g/L 1.308 D

NaCN Consumption kg/batch 25 E

NaOH Consumption kg/batch 2.5 E



17.2.6 Pre-Oxidation, Leach and Adsorption Circuit

The overflow from the cyclone overflow will pass across to the leach train area. The pulp will

pass through a Delkor trash screen which removes, in particular, wood pulp which would

choke the carbon screens in down-stream processing. The trash falls into the tailings pump

hopper and is pumped with tailings out to the final tailings dam.

The pulp (now without trash) falls into the pre-oxidation/leach feed pump hopper from whence

it is pumped to the first pre-oxidation tank at the opposite end of the leach train. An automatic

sampling system will be installed in this line so that a representative sample is retained for

accounting purposes. The leach circuit is divided into a pre-oxidation stage of 12 hours

followed by a CIL circuit. The pre-oxidation stage consists of three large mechanically agitated

tanks with a useful capacity of approximately 2,076 m3 each. Agitation will be provided by

dual rubber coated impellor agitators with axial flow or similar. These tanks will be oxygenated

using a PSA oxygen supply system. Oxygen will be added by sparges inserted into the exit

side of recirculating pumps which will take pulp from the lower part of each tank and pump

back into the body of the tank. It is intended to maintain oxygen contents in the order of

20 ppm in liquid. This oxidation process is necessary to ensure that the tellurium is oxidized

and therefore ceases to have a passivating effect on the gold. There will be a height

difference of 150 mm between these pre-oxidation tanks. But the platform over these three

tanks will be at the one level. There will be a facility which allows any of the tanks to be by-

passed.

The pulp will overflow from the last pre-oxidation tank into the first of a series of 6 CIL

mechanically agitated tanks each of volume of 2076 cubic metres. This allows for a leaching

time of 24 hours. There will be no height difference between successive tanks in this case and

the level platform over the tanks will be used to support the agitators as well as all other

ancillary equipment. The leeway between the top of one tank and the level of the overflow will

be one metre. The pulp will flow from one tank to the other through a rotating Kemix

interstage screens which has a pumping capacity to maintain a head and thus no height

difference is required. The freeboard allows for variations in pulp density from one tank to

another which affects the performance of the Kemix interstage pump/screens.

New or regenerated carbon will be added to the last CIL tank at a rate to maintain the carbon

contents of the last and first CIL tanks at 15 g/L, while the other four tanks will be maintained

at 7 g/L. When this addition of carbon is made using regenerated carbon, the carbon will be

passed over a vibrating screen to ensure that any carbon less than 1000 µm is rejected and

sent directly to tails, rather than re-entering the circuit.

Reactivated carbon will be provided at 6 x 12 mesh sizing. Carbon will be transferred up-

stream, against the flow of pulp, using Krebs Millmax submersed pumps (or equivalent) with

recessed impellors. The transfer system will include provisions to pump to one of two CIL

tanks by the use of valves.



Loaded carbon from the first CIL tank, at appropriate intervals, will be pumped over a loaded

carbon screen, the undersize pulp returning to circuit whilst the oversized carbon (+800 µm),

after thorough washing with sprays on the screen, falls into an appropriate conical bottom acid

proofed fibre-glass storage vessel.

Sodium cyanide solution will be added to the first, second and third CIL tanks at a rate to

maintain the exit free cyanide concentration at approximately 50 ppm. The rate of addition will

be controlled by a continuous cyanide analyser which will continuously sample cyanide

content of the first, the last CIL tanks as well as the product of the cyanide destruction tank.

All the pre-oxidation and CIL tanks will be encircled by a bund with capacity to include the

total contents of the biggest tank in case of total failure or if any tank needs to be emptied for

any reason.

An overview of the pre-oxidation, leaching and detoxification is presented in Table 17.2.6_1.

Table 17.2.6_1

Mara Rosa Project Data Sources for Pre-oxidation, Leaching and Detoxification


Nominal Recovery Data

Solids Feed Rate tph 317 C

Feed Pulp Density % w/w 44.1 D

Feed Flow Rate m3/hr 518 G

Design Mill Feed Grade g/t 1.46 D

Leach Feed Gold g/hr 463 D

Gold Lost in Solid Residue g/t 0.12 T

Gold Lost in Solid residue g/hr 39 D

Gold Lost in Solution Tails g/t 0.01 F

Gold Lost in Solution Tails g/hr 4 D

Gold Recovered g/hr 420 D

Global Gold Recovery % 92 D

Trash Screen

Type Delkor, size 20 m2 E

Screen Feed Rate m3/hr 519 C

Screen Aperture microns 800 E

Spray Water m3/hr 15 E

Leach Feed Pumps

Type Rubber Lined Horizontal Centrifugal

G

Size Warman 8 x 6 or similar E

Feed Flow Rate m3/hr 518 D

Material Pulp, maximum particle size 0.8 mms

F

P80 45 µm, pulp density 44% w/w

Pressure kpa 200 F


Leach Feed Sampling System Primary and Secondary Timed Sampling System




Pre-Oxidation Tanks

Number of Stages 3 G

Total Residence Time hr 12 T

Effective Volume of each m3 2076 D

Method of Oxygen Production PSA (Pressure Swing Adsorption)

G

Method of Oxygen Addition Sparges in Recirculating Pumps

B

Target DO Content ppm 20 G

CIL Tanks



Effective Volume of Each tank m3 2076 D

Total residence Time hr 24 T

Minimum Carbon concentration g/L 7 G

Maximum Carbon Concentration g/L 15 G

Total Carbon in Circuit t 120 D

Design Carbon loading g/t 2710 D

Average carbon Advance tpd 4 B

Carbon Advance Period hr 8 G

Carbon and Pulp Forwarding Rate m3/hr 35 D

Carbon Advance Method Recessed Impellor Pump G

Loaded carbon Screen

Type Horizontal vibrating with 0.8mm G

polyurethane screens

Size m 1x2 E

Design Feed Rate m3/hr 50 G

Intertank Carbon Screens

Type Cylindrical Mechanically Swept G

Screen Material Stainless Wedge Wire G

Screen Aperture mms 1.0 G

Safety Carbon Screen

Type Horizontal vibrating with 0.8mm screens

G

Carbon Sizing Screen

Type Horizontal vibrating with 1.0 mm screens

G

Thickener

Diameter m 30 A

(using 0.5 t /m2 per hour settling rate)

Thickener Underflow Pumps

Type Horizontal Rubber-lined centrifugal

G

Size 8 x 6 E

Flow m3/hr 328 D

Material Pulp with P80 45 µm F

Density % w/w 60 G

Pressure kpa 100 A

Detoxification Tanks


Hours per Stage 2 G




Effective Volume of Each tank m3 656 D

Tailings Pumps

Type Rubber Lined Horizontal Centrifugal

G

Size 8 x 6 E

Flow m3/hr 328 D

Material Pulp, 60% w/w solids P80 45 µm

F

Pressure kpa 500 F

Tailings Sampling System Primary and Secondary Timed sampling System

17.2.7 Carbon Elution Circuit

The loaded carbon arisings will be about 4 tpd assaying up to 2000 g/t of contained gold will

be water washed and then acid washed by recirculating a 3% hydrochloric acid solution

through the carbon acid wash column. After acid washing, the contaminated acid solution

(now with calcium salts amongst others) is rejected via the plant tailings. The carbon will then

be transferred to the strip pressure column. Carbon is eluted by passing a hot solution

containing 2.0% caustic soda and 0.2% cyanide under pressure at the minimum rate of

1.5 bed volumes per hour. In the intended system the subsequent gold bearing solution

passes through an electrolytic cell where a current of about 1000 amps/m2 is applied at a

voltage of ±4.5 volts before being recycled back to the elution vessel. This recycling,

commonly known as the Zadra system of elution, continues for about 12 hours until the gold

content in the carbon is approximately 50 ppm. At this point the circuit is cooled down and the

carbon washed and transferred to a holding hopper above the regeneration furnace. Carbon

loses its activity gradually unless regularly regenerated by heating up to 600 °C in a controlled

atmosphere.

The strip solution (now depleted of gold and somewhat diluted) is returned to its holding tank

where further cyanide/caustic is added before a next stripping sequence. This solution has to

be totally renewed at intervals (every third or fourth strip) when old solutions are returned to

the leach circuit.

The cells are composed of stainless steel wool cathodes supported by a cage-like

arrangement and stainless steel anodes and will be cleaned out at least once a week. Gold

from the cell floor and from the cleaning of each cathode using high pressure sprays will be

dried, calcined and then smelted in a small diesel fired furnace before being sampled,

weighed and transported to the refiner.

The carbon after regeneration is returned to the last CIL tank via a sizing screen to ensure

that finer carbon less than 1,000 µm does not get returned to the CIL circuit but is sent to tails.

An overview of the elution is presented in Table 17.2.7_1.



Table 17.2.7_1

Mara Rosa Project

Data Sources for Elution


Elution Type Pressurized Zadra E

Required Elutions per Week 7 E

Elution Column Capacity t 4 E

Carbon Bulk Density t/m3 0.48 A

Elution Solution Strength % w/w NaCN 0.2 E

Elution Solution Strength % w/w NaOH 2.0 E

Elution Flow Rate, (Bed Volumes) BV/h 2 E

Elution Flow Rate m3/h 20 E

Volume of Elution Tank m3 20 E

Fresh Eluate Make-Up No. of Strips 4 E

Elution Time hr 12 E

Elution Temperature centigrade 140 E

Elution Operating Pressure kpa 600 E

Elution Heater Type LPG E

Number of Cells 2 G

Cell configuration Parallel G

Cell Size mms 600 x 600 E

Number of cathodes per Cell 10 E

Cathode Type Stainless Steel Mesh 0.156mm E

Number of Rectifiers 1000A 2 G

Gravity Electrowinning

Number of Cells 1 G

Cell Size mms 600 x 600 E

Number of Rectifiers 1000A 1 G

Carbon Regeneration

Carbon Feed Hopper Capacity m3 11.2 E

Reactivation Kiln Type Rotary, horizontal E

Feed Rate of Carbon kg/hr 200

Operating Temperature centigrade 650-750 E

Feeder Type Screw Feeder E

Fuel LPG E

Quench Tank Capacity t 4 E

Gold Room

Preparation Sludge Cone settling, Drying and Calcine Ovens

E

Smelting Pendant Controlled LPG Barring Furnace

E

Exhaust Systems Extraction Hoods and Fans for cells and furnaces

E

17.2.8 Tailings Disposal

The last CIL tank overflows to a vibrating screen which acts a carbon safety screen with

apertures of 0.8 mm. Usually very little carbon is collected by this system. If the volume and



grade permits this can be sold to third parties who treat the carbon to scavenge out any gold

and reclaim the activated carbon (now fine) for other uses (colour removal of liquids, etc.)

The underflow from the safety screen goes to a thickener whose purpose is to recover and

return some of the high pH solution containing some cyanide solution back to the milling

section at the same time reducing the quantity of cyanide to be subsequently destroyed. The

thickener underflow at around 60% solids will be pumped to the first of two cyanide

destruction vessels, 655 m3 each, with a residence time of two hours in each, before

overflowing to the tailings pump box. Destruction will be carried out by a mixture of sodium

metabisulphite, air and copper sulphate. A second automatic sampling system is required for

accountability of the leaching circuit and this can either be in-line or as a “cross-flow cut” from

the flow entering the tailings pump hopper.

The pulp pumped out by the tailings pumps will be the only outlet of plant rejects. All other

plant streams are recycled. The tailings line will consist of a suitably sized high density

polyethylene pipe with thickness appropriate to the applied pressure. There will be flow

meters/ pressure indicators, etc., on this pipe-line. These tailings will be the only final residue

from the processing plant. All sump pumps throughout the plant will recirculate flows back to

the circuit.

The tailings dam will consist of a large area with sides raised using mine waste lined with a

compacted clay-type material to reduce seepage. The pulp will be allowed to settle out.

Decantation towers will be placed within the tailings pond in order to reclaim water. This water

will be used as process water being pumped to the plant via a water reservoir which will also

receive make-up water from nearby rivers in the rainy season. In a normal year there is a

considerable deficit in the water balance, thus the reason for the insertion of the water

reservoir whose level should be gauged as to be full at the beginning of the dry season,

(April).

17.2.9 Reagents

As already described lime will be delivered in tankers for pneumatic distribution to the lime

silos.

The road transport of sodium cyanide is a highly controlled manoeuvre which only certain

transport companies have permission to transport. Sodium cyanide will be transported as

solid briquettes in thick plastic one tonne big bags inside wooden crates. When required a

fork-lift will transport the boxes and the plastic big bags are lifted into a ventilated container

with appropriate ventilation and the bag pierced in such a way that the solids fall into an

agitated tank with a capacity of eight cubic metres. The dissolved cyanide solution is then

pumped to a similarly sized stock tank for recirculation to the plant. A pump will pump the

liquid (at around 20% cyanide) through a manifold from which cyanide is monitored using

meters into any of the first three CIL tanks.



Caustic soda (sodium hydroxide) will be received as a solid and added directly as required

into a five cubic metre strongly agitated stock tank. This tank will be replenished daily by the

addition of the appropriate quantity of make-up water followed by the solid addition. The final

concentration will be in the order of 200 g/L.

Both the caustic and cyanide preparation systems will enjoy a common bunded area as these

products are compatible and have an appropriate sump pump in case of spillage.

Hydrochloric acid is received in tankers containing acid of normal commercial concentrations

(about 33% HCl) and stocked in a fibre glass tank containing a total of 30 m3 contained within

its own separate bund and with its own sump pump. This tank will have its own bund area

with appropriate sump pump.

Copper sulphate and Sodium Metabisulphite will be received as solids and will have to be

dissolved in a similar way as the caustic soda by adding directly to an agitated tank which

also acts as a storage vessel. These two tanks will be placed in a common bund area with a

sump pump.

Flocculent will be prepared by the usual eduction systems whereby the solids are dissolved

and kept in a storage vessel with pumps to add flocculent to the feed to the thickener.

An overview of the reagents is presented in Table 17.2.9_1.

Table 17.1.3_1

Mara Rosa Project



Mill Balls

Primary Mill Type 70 mms E

Consumption grams/tonne of ore 500 F

Daily Consumption t 3.4 D

Annual Consumption t 1250 D

Secondary Mill Type 25 mms E


Daily Consumption t 3.4 D


Cyanide

Consumption, leaching kg/tonne of ore 0.26 T

Elution and Intensive Leaching kg/day 65 E

Average Total Consumption kg/day 2044 D

Monthly Consumption t/month 61 D

Make-Up Concentration g/L 200 G

Size of Daily Make-up Batch m3 10.2 D

Weight of Unit Delivery t 25 E

Frequency of Deliveries average every 12 days E

Annual usage t 736 D




Sodium Hydroxide

Consumption, (elution) kg/day 400 E

Consumption, (gravity) Kg/day 2.5 E

Make-up Concentration g/L 200 G

Size of Mate-Up Batch (6 days) m3 12.1 G

Total Consumption t/month 12.1 D

Weight of Unit delivery t 25 E

Frequency of Deliveries per annum 6 D

Total Annual Consumption 147 D

Oxygen

Pre-Oxidation kg/tonne of ore 0.8 F

Daily Consumption t 6 D

PSA Oxygen Plant 6 tons per day D

Lime

Design Consumption kg/tonne of ore 2.1 T

Daily Maximum Usage t/day 15.2 D

Addition Point Mill Feed Conveyor G

Annual Usage t 5250 D

Pre-Leach pH level pH 12 T

Number of Silos 1 E

Silo Capacity t 60 C

Truck Capacity t 25 E

Delivery Period 2 days in three E

Hydrochloric Acid

Supply Strength %w/w 33 E

Solution, specific gravity t/m3 1.15 E

Consumption m3/strip 0.8 G

Consumption m3/month 24 D

Consumption m3/annum 285 D

Size of Container m3 12 E

Storage capacity m3 30 G

Delivery Frequency Every two weeks D

Activated Carbon

Consumption grams/tonne of ore 50 G

Consumption, average kg/day 342 D

Consumption, average t/annum 125 D

Delivery Method Big Bags of approximately 550 kg

E

Initial Fill t 120 D

Initial Order t 130 D

Subsequent Orders Monthly, 11 tonne batches E

Sodium Metabisulphite



Make-Up concentration g/L 200 G

Solution Consumption m3/day 8.6 D


Delivery Size, solid in 20 kg packets

t 25 E

Delivery Frequency Every two weeks D

Copper Sulphate






Make-up concentration g/L 200 G

Solution Consumption m3/day 3.43 D

Consumption t/annum 250 D


25 E

Delivery Frequency Every month D

Flocculant



Consumption t/annum 25 D


t 5 E

Delivery Frequency Every second month D

17.2.10 Water Distribution

Water is essential for the processing plant. After initial start-up there will be a continuous

source of water returning from the tailings dam as the pumps in the decant towers are

operated. To add flexibility and also to avoid excessive tailing dam height to retain rain water

during the rainy season, a water storage reservoir with approximate volume of 500,000 m3 will

be built which will receive water from the decant towers as well as river water from a pump

located near to a local river which will take off sufficient water during the rainy season to

ensure that there will be water available throughout the dry season. There will also be water

from the mining pit: (whilst mining, there will be ingresses of water to the main pit and this is

required to be pumped out to maintain the floor of the pit workable).

The plant will operate under the system of zero discharge to the environment.

Water from the water storage reservoir will be continually pumped to the process water tank

of volume approximately 1,000 m3. This source of water will also be used (together with water

from the pit) to dilute thickener feed which operates more efficiently with feeds approximating

to 20% solids. The overflow from the thickener will in turn recirculate to feed the process

water tank which will consist of a steel tank. Pumps will distribute process water around the

plant and includes an alternative route to fill or compliment the water addition to the thickener

when necessary.

The maximum requirements of water will be 5,000 m3/d. Any water supply deficiency (other

than that returning from the tailings dam) will be made up from wells or from surface water

streams. Good quality water can be sourced continuously from wells in the area.

In addition to the process water tank there will be a raw water tank. This will be fed good

quality water from surface water streams or from the water storage facility. This will supply

water in some discreet areas in the plant such as carbon wash, seal water pumps, reagent

make-up, emergency showers, feed to the potable water plant, etc. A smaller tank of 250 m3,

also of steel, is considered adequate. This will also be the feed to the fire system which will



consist of a pressurized manifold piping system with jockey pump. This tank will be kept full

with a continuous overflow to the process water tank.

Finally a potable water plant will produce potable water from local well water which will be

stored in a 50 m3 steel tank for distribution around the plant.

An overview of the water consumption is presented in Table 17.2.7_1.

Table 17.1.3_1

Mara Rosa Project



Raw Water

Raw water for potable water

m3/day 96 F

Loaded carbon Screen m3/day 120 G

Carbon Transfer m3/day 60 G

Acid Wash m3/day 5 G

Elution Circuit, average m3/day 15 F

Cyanide Preparation m3/day 10 F

Caustic Preparation m3/day 5 F

Sodium Bisulphite Preparation

m3/day 20 F

Copper Sulphate Preparation

m3/day 5 F

Seal water m3/day 425 F

Road Wetting m3/hr 264 G

Total Fresh Water Usage m3/day 1025 D

Total Fresh Water Usage m3/hr 43 D

Process Water

Milling m3/hr 374 D

Trash Screen m3/hr 15 E

Safety Screen m3/hr 10 G

Raw Water re-entering system

m3/hr -28 G

Return from Thickener m3/hr -300 D

Total Process Water Usage

m3/hr 71 D

The plant is in deficit for water when the water tied up in the tailings is considered. Water

make-up is a combination of local wells (better quality water), water pumped from a local river

in the rainy season as well as contributions from pit dewatering.

17.2.11 Electrical Power Supply

A 138 kV power line will be constructed from Porangatu for a distance of 63 km. A sub-station

near the mills will monitor the incoming energy and with appropriate rectifiers will reduce this

voltage to approximately 6,600 kV. This supply will be used to power the mills. This same



voltage will be used to distribute the power to further sub-stations where the voltage will be

reduced to the standard three phase 460 v for use at the smaller motors.

There will be an diesel generating set for emergency power. This will be of sufficient capacity

so as to power all the pre-oxidation, CIL agitators and cyanide destruction tank agitators,

maintain all lighting and some emergency services.

17.2.12 Laboratory

The laboratory will be of the traditional fire assay type with finish by solvent extraction using

MIBC followed by Atomic Absorption. There will be sufficient capacity to treat the mine

samples as well as the daily process samples (estimated at an average total of 100 per day of

various types). The laboratory will include capacity to crush and pulverize solid rock samples

with associated exhaust systems. As well as chemical analyses, regular size analysis of

different streams will be carried out. Sample preparation will be carried out 24 h/d with results

being published on a daily basis by the day shift technical staff.

17.2.13 Plant Security

The gold-room with its attendant operations (electrowinning, gold sludge retrieval, calcining

and barring furnace) will be in a separate C container and will include the supply of a gold-

room security system comprising PIR alarm system with keycode access, cell PIR’s, safe

vibration sensor, four off Samsung surface mounted colour cameras, quad splitter,

17” observation LCD monitor and digital recorder system with Ethernet rebroadcast capability.

In addition the plant entrances will ensure that all visitors and employees will be vetted and

properly identified when entering the plant area. It is intended that a consultant be employed

before construction to enable the question of security to be addressed adequately.

17.2.14 Flowsheets and Layout

Figures 17.2.14_1, 17.2.14_2 and 17.2.14_3 show the basic flowsheet block diagram,

process flowsheet and plant layout for the proposed plant design.



Figure 17.2.14_1

Plant Flowsheet Block Diagram

Britagem Primário

Oversize

Undersize

19mms

OVERSIZE

LEACHED LIME ADDITION

CONCENTRATE

MILL DISCHARGE

TAILS

CONCENTRATE CATHODE

Intensive Leach Reactor

UNDERSIZE SODIUM STRIPPED

P80 of 45 microns CYANIDE CARBON CATHODE

Bombas p/ Alimentar Ciclones REGENERATED CARBON

PREGNANT SOLUTION

LOADED WASHED

CARBON CARBON

DILUTE HYDROCHLORIC ACID SOLUTION WITH SODIUM HYDROXIDE

AND CYANIDE

OXYGEN

PLANT

TAILINGS

SODIUM METABISULPHITE

AND COPPER SULPHATE

WITH AIR

ORE PRIMARYCRUSHER

SECONDARYCRUSHER

VIBRATINGSCREEN

TERTIARY CRUSHER

SiloPRIMARYMILL

SECONDARYMILLCYCLONE FEED

PUMPS

CYCLONES

GRAVITYCENTRIFUGE

INTENSIVELEACH REACTOR

ELECTROLYSIS

PRE‐OXIDATIONTANKS

CIL

THICKENER

ACIDWASH ELUTIONCOLUMN

CARBONREGENERATION

ELECTROLYSIS

CALCINATION

SMELTINGFURNACE

GOLD BULLION

DETOX TSF



Figure 17.2.14_2

Plant Process Flowsheet



Figure 17.2.14_3

Plant Layout



17.2.15 Main Process Equipments

The main process equipment is listed in Table 17.2.15_1.

Table 17.1.3_1

Mara Rosa Project


No in operation

No installed

Total flow Unit flow Total flow Unit flow

solids t/hr m3/h t/h each m3/h each

Crushing Circuit 0 0

ROM Concrete Dump Pad 1 1 381 211 381 211

Vibrating feeder with Grizzly 1 1 381 211 381 211

Jaw Crusher 1 1 381 147 381 147

Primary Screen 1 1 381 254 381 254

Secondary Cone Crusher 1 1 343 247 343 247

Secondary Screen 1 1 514 370 514 370

Tertiary Cone Crusher 1 1 171 123 171 123

Milling Circuit

Primary Ball Mill 1 1 317 222 317 222

Secondary Ball Mill 1 1 634 444 634 444

Scyclone Feed Pumps 1 2 951 931 951.00 931.00

Hydrocyclone Nest 1 1 951 931 951.00 931.00

Gravity Circuit

Centrifugal Concentrator 1 1 200 201 200.00 201.00

Intensive Leach Unit 1 1 200 201 200 201

Pre-Oxidation and leach Circuit

Linea Screen 1 1 317 504 317.00 504.00

Pre-Oxidation Tanks 3 3 317 519 317.00 519.00

Pre-Oxidation Agitators 3 3 317 519 317.00 519.00

CIL tanks 6 6 317 519 317.00 519.00

CIL Agitators 6 6 317 519 317.00 519.00

Thickener 1 1 317 723 317 723

Detox Unit

Elution Circuit

Elution Package

17.2.16 Automation and Control

Introduction

The purpose of this document is to define the overall operating control philosophy of the plant.

Control rooms with visual observation of the processes will be situated in the Crushing,

(monitoring only) and Milling areas. Overall control will consist of distributed control systems



with a series of local programmable logic controllers (PLC) providing control at field level.

There will be two separate control rooms one adjacent to the primary crusher and controlling

the crusher circuit up to the fine ore storage bin and the other situated on the mill platform

which will control all the rest of the operations starting from this same silo up to the final

tailings pumps. The status of each equipment can be monitored on PC based Supervisory

Control and Data Acquisitioning (SCADA) screens which will show a series of mimic screens

representing various areas of the plant and will show parameters such as flow, level,

temperature, pressure, status and set-points.

The plant is designed to allow the operator to monitor the whole plant including water services

and control all important aspects of the plant. Motor starters, main isolators and distribution

board feeders are located in the Motor Control Centres (MCC). Each MCC also contains a

PLC which, along with the SCADA system makes up the Process Control System (PCS). The

PCS provides the interface between drives and instrumentation, and the operators. Hardwired

outputs and inputs for plant equipment and field devices are interface through these PLC’s in

which plant start up and shutdown sequences and interlocks are programmed.

The majority of the drives will be remotely operable from the PCS. Most electrical drives can

be operated remotely from the control rooms or locally from control field stations.

The mode of operation is selected by the PCS operator. When “Manual “mode is selected the

drive can be remotely operated from the control room. When “Auto” is selected the drive can

only be operated by the sequencing logic. For these drives to operate all associated

permissive and process interlocks are required to be in a healthy state.

To operate the drive in Local, the PCS operator must be requested to select “Local” mode for

the requested drive. The drive can then be operated in the field by use of local start and stop

buttons. When operating in this mode, permissive and processing interlocks are by-passed,

(this is also a requirement for maintenance purposes) and therefore the appropriate safety

measures (such as lock out of adjacent equipment) must be carried out.

The field mounted local control stations (LCS) will include a “START” pushbutton and a

latching mushroom-head “STOP” button. The stop buttons are hardwired to stop the drive

regardless of the mode of operation.

Every major equipment, mill, crushers, etc. plus in some cases a whole section (crushing

circuit), have start-up horns and flashing lights.

Crushing Circuit

As in all major sections in the plant the start sequence includes the initiation of flashing lights

and horns. All the three crushers are stopped and started independently of the crushing circuit

sequence. Also the fact that there are two intermediary stockpiles means that each crusher

can be run independently. If the appropriate crushers are on line then the start-up sequence

for each one can be initiated providing that the drives are in Auto and continues with the start



up of the appropriate screen and conveyors (in the case of the cone crushers) whilst the jaw

crusher requires the two product conveyors 100-CV-001 and 002 before initiation of the

vibrating feeder 100-VB-001. Again in the case of the cone crushers the appropriate screens

and conveyors can only be switched on when the respective crusher is functioning and only

when all relevant conveyors and screens are functioning will the vibrating feeders be able to

initiate production.

The operator will control the vibrating feeder/grizzly velocity to require the desired tonnage

using the weightometer 100-WE-001 on the primary screen conveyor. Similarly the operator

can adjust the rate of withdrawal from the two stockpiles by altering the frequency of the

respective vibrating feeders.

The PCS operator will not be required to individually start each motor, (other than those

mentioned above) by selecting the sequence in automatic the electric motors will themselves

start up in the designated order with appropriate small time intervals between each equipment

start. If any equipment failure occurs during the start-up the sequence will discontinue.

To stop the crushing circuit, there will also be an appropriate stopping sequence(s). Normally

the circuit will be crushed out so it will be programmed that the various equipments cease at

different time intervals. The presence of two intermediate stockpiles will render this process

easier.

If the down-time is expected to be prolonged then all three crushers will be turned off.

If a failure occurs then it will be programmed that the relevant part of the crushing circuit will

also immediately stop. Thus, for instance, in the case of a failure of the final conveyor belt

feeding the fine ore bin, 100-CV-009, only the jaw crusher and associated equipment will

continue functioning feeding the first intermediate stockpile. Failures of conveyors occur

principally because of motor or actual belt failures (or abnormalities) and are sensed by

velocity sensoring devices as well as positioning sensors and belt tear sensors on each belt.

Ball Mill Feed System

The ball mill feed system consists of various feed conveyors drawing crushed ore from

underneath the fine ore bin. These feed a final mill feed conveyor, 100-CV-011. This will be

initiated independently when the both ball mills status are OK and running. By manipulation of

the number and velocity of the belt feeders the operator can maintain the desired feed rate as

registered by the weightometer 200-WE-01. Again sirens and flashing lights are part of the

start sequence and warn when the conveyors are about to run. The screw feeder on the lime

silo can only operate if the final mill feed conveyor is functioning and will automatically switch

itself off when this conveyor is not operating.

Preparation for Starting the Ball Mills

Before either ball mill can be run the cyclone feed pump has to be running which in turn

means that the leach section trash screen, pre-oxidation/leach feed pump, carbon safety



screens, thickener discharge pumps, (Detox feed), and tailing pumps have all to be running.

Thus correct procedure involves starting the tailings pumps (after ensuring that the gland

pumps are supplying the correct pressure). This is an independent operator order though of

course there are pre-conditions involved such as if the tailings are apt to receive reject, the

tailings line is functioning, whether there is water being added into the pump hopper (the

pump cannot run empty), etc. Then the same applies to the thickener discharge pumps which

feed the Detox and the pre-oxidation/leach feed pumps, the trash screen and the carbon

safety screen. The last item to be started before the mill itself is the cyclone pump. The same

philosophy applies here, the operator consciously starts the appropriate pump providing the

conditions are suitable (gland water pressure adequate, appropriate valves open and

sufficient cyclones on line).

There will be two methods of operating the cyclones. In the one case the velocity of the

cyclone pump will be controlled by the level of pulp in the pump hopper. A level sensor will

measure the height of pulp in the pump box and the pump speed will be adjusted to maintain

this level constant. There will also be a pressure sensor on the cyclone distribution box and

the facility to control the pump velocity to maintain a constant pressure. This is the preferred

method of operating the cyclones but means that the water addition to the cyclone hopper has

to be varied using a flow control valve which in turn is controlled by the level sensor to

maintain the level of pulp constant. This also involves increased attention with respect to the

number of cyclones on line which is why some of the feed valves on the cyclones will be able

to be controlled automatically by the operator. The control room operator will be able to

choose which system to operate under or indeed operate the system in manual mode

whereby he selects the number of cyclones, the velocity of the pump and the water addition

rate; this will be especially relevant during start-up and shut-down when flows tend to fluctuate

greatly.

Ball Mills

Each ball mill will have its own start sequence and will depend on a large number of status

signals. For the sequence to activate, the status of not only the above preparatory equipments

(pumps, screens, etc) must be shown to be running, but also a series of permissive signals

must be present. These include, trunnion high and low pressure pumps running, pinion

bearing pump running, trunnion and pinion bearing temperatures considered healthy,

lubrication flow and pressure switches healthy, lubrication oil cooling fans ready and the girth

gear lubrication system in operation.

Thus when the sequence for starting the mill is initiated the above oil pumps will start

operation in the correct order until finally the mill motor will initiate. Then if there are no

negative signals from the various sensors and meters, a request to engage the clutch will

appear and when actuated, the mill will begin to turn.

It should be noted that if a ball mill has been down for an extended period (to reline for

example) then the mill must be inched prior to starting. This is a device whereby a small motor

with a large mechanical advantage is used to revolve the mill very slowly so that the internal



contents are loosened up. Note that the trunnion lubrication pumps should be in operation

during this procedure. When inching, the mill motor must be correctly locked out.

The water addition to the mill feed will be added by applying a ratio to the weightometer 200-

WE-001 sensing the feed rate on the mill feed conveyor 100-CV-011 and thereby controlling a

control valve in the water addition line.

Lime addition will be added onto the mill feed belt. The monitoring device for pH in the first

pre-oxidation tank will display a value in the control room and the rate addition will be

controlled by the operator by increasing or decreasing the velocity of the screw feeder.

Gravity Circuit.

This consists of a pump which takes a fraction of the ball mill discharge and pumps it over a

vibrating screen where the undersize is fed to a Knelson or Falcon Centrifuge. Control is

manual from the control room. There is no sequencing. However the Knelson or Falcon itself

automatically follows a fixed routine dictated to it from a control panel located in the Acacia or

Gekko unit. This control panel is part of the gravity concentrate package and determines the

parameters by which the processes are controlled such as interval between washing out of

the centrifuge, quantities of reagents use in the intensive leaching process, etc. The control

room operator has no control of any of these parameters confining himself to starting and

stopping the gravity feed pump, the vibrating screen and the gravity concentrator. The

intensive cyanidation sequence is divided into sections each of which require initiation on a

local control panel. These include discharge of solution after electrowinning (after analysis of

a solution sample), transfer of the next batch of pregnant solution, initiation of the concentrate

leaching sequence and finally discharge of the concentrate residue.

Leach and Detox and Tailings Circuit

The PCS will monitor the status of all the principal drives in the pre-oxidation, leach and

adsorption section. There will be level sensors in all the pump hoppers and the variable speed

pump motors will be adjusted to maintain a constant level. This applies to leach feed pumps,

the thickener discharge pumps which transfer reject to the Detox unit as well as the final

tailings pumps.

Shift samples of the feed to the pre-oxidation/leach circuit will be taken by an automatic

sampling device. This will be located on the outlet of the pre-oxidation/leach feed pumps and

produce a sample bucket of pulp for the operators to take to the plant laboratory. Size

analyses and gold grades will be measured on a regular basis. Similarly automatic sampling

of the final reject before it gets pumped to the tailings dam will produce shift samples which

will also be assayed for gold, cyanide content and size by size fractions measured.

A monitoring device for cyanide concentration will be operating which will monitor cyanide

content in the first and last tanks in the leach circuit as well as in the final detox tank. However

cyanide additions will be controlled by the control room operator who raises or lowers the



cyanide solution rate according to the value given by the monitoring device (which is checked

twice each shift using normal titration: pH is checked in the same manner).

Carbon concentrations in the tanks will be carried out by the leach operator on a shift basis

using a manual screening device and oxygen contents will also be adjusted locally using

valves and flowmeters. Oxygen contents in the pulp will also be determined locally by using

handheld oxygen measuring devices.

Transfer of the carbon from one tank to the other as well as transfer of loaded carbon from the

first CIL tank will be carried out on a manual basis by the PCS operator with coordination from

the leach operator. A routine will normally be followed in order to maintain correct carbon

concentrations in each CIL tank.

Elution

This whole section will be subject to a control system supplied as part of the elution and

carbon regeneration package supplied by Como Engineering. The system is fully automated

PLC controlled, for all the major sequence steps from acid washing through to eluting, barren

carbon washing, regeneration transfer, regeneration and carbon quenching and carbon

transfer back to the CIL circuit and will communicate with the main PCS.



18 PROJECT INFRASTRUCTURE

18.1 General Criteria Adopted

The project infrastructure has been scaled as appropriate to the mine production rate

described in Section 16. The specific industrial facilities to service the mining area, the

manufacture of explosives and storerooms, supply of mining equipment and mechanical

repairs were studied by Amarillo Gold / ONIX.

The asphalted road from Mara Rosa to Amaralina fortunately by-passes the town centre of

Mara Rosa thus avoiding inconveniences to the towns people with respect to excess traffic

through the town or excess noise at night. A good serviceable dirt road will be implanted from

this main asphalted road to the plant area. The car park for visitors and employees will be

outside the fenced area and outside the central offices where the administration and

metallurgy offices will be located. Adjacent to the offices will be the changing rooms and the

main canteen.

A fenced area around all the industrial buildings including the electrical sub-stations will

include a reinforced gate with an attached guard post.

The general layout of the mine site is presented in Figure 18.1_1 and items included in the

industrial infrastructure of the Mara Rosa Project are described in the following sub-sections.



Figure 16.2.4_3

Master Plan – Site Layour



18.2 Procurement and Distribution of Water

18.2.1 Types of water

The types of water will be used within the mine and plant areas as follows:

Raw water;

Drinking water (derived from the raw water) and

Process water (consisting of water recovered in the tailings thickener and tailings dam).

The reclaimed water in the tailings pond will be composed of water supplied by the

detoxification of the tailings slurry and rain water (from the catchment of the surface of the

tailings dam). The positive water balance of the tailings pond results from the rainfall. The

overflow of the tailings dam will occur as a result of the contribution of these waters. The

water in the tailings dam overflow will not necessarily be continuous. The "water level" will

vary according to season, but without risk of running dry.

18.2.2 Raw water

The project area does not host favourable groundwater potential with rain runoff being the

only alternative source in the region.

In addition to recirculated water, the plant will need new water supply, which will be captured

at the clean water dam. Water will be pumped from sumps in the mine pits, reclaimed water

will be pumped from the tailings pond, and rainwater will be captured in the areas of influence

of this basin.

The clean water dam will have a length of 360 m, height of 12 m, and storage capacity

estimated at 514,000 m³.

Alternatives are being studied which include capture of water from the river “Rio de Ouro”

(literally Gold River) during the wet season. This is 1.5 km from the plant site, from which

pumping to a storage tank at the processing plant will be undertaken.

18.2.3 Drinking Water

Approximately 400 employees may be on site at any time, in the mine, the plant, maintenance

workshops and administrative offices. Consumption is estimated at 5 m3/h and an installed

capacity of 15 m3/h is provided for in this study.

Drinking water will be used for use in cafeterias, water fountains, general health and

emergency showers, eye wash, that will be implemented in the laboratories, preparation of

reagents and wards.



18.2.4 Process Water

Process water will be sourced from reclaimed water in the dam and that recovered in the

tailings thickener. Each of these sources will have a pumping station to deliver water to the

process water reservoir.

The uptake of clean water in the dam will be made with a vertical pump system positioned on

a floating raft on the dam, and pumping to the reservoir.

The reservoir will supply the process water by gravity, water supply for the mills, dilution water

pump boxes and screening Knelson concentrator, leaching and CIL tank, the "underflow "

cyclones and cleaning of the "underflow" pipe of the tailings thickener.

For water supply to screens, the tank leaching circuit (CIL), and load lines for the preparation

of reagents it will be necessary to apply a "booster" pump to provide process water at high

pressure.

18.3 System and Distribution of Electricity

18.3.1 Supply of Industrial Units to Mara Rosa

The Mara Rosa Project requires an estimated power supply of 12.7 MW. This will be supplied

by the power utility CELG D (Central Goiás Electrical Distribution SA) through a 138 kV

transmission line of 64 km derived from the substation in the city of Porangatu.

The plant's main substation will be constructed with four substations equipped with

transformers, power frames, CCMS and all other electrical equipment and automation.

There will be a substation close to the mine pit to supply power to the filling station, mine

workshop, mine office and dewatering pumping system.

In order to minimize additional environmental impacts such as deforestation and right of way,

the route of the transmission line should follow to the fullest extent possible, the route of the

paved road to be built between the city of Porangatu and the project .

The cable to be used shall be of the optical fiber (OPGW), specially designed for installation in

overhead power lines.

18.3.2 Distribution of Energy within Project Area

The power distribution system for the different load centers and facilities, shall be of the

simple overhead distribution networks, such as "isolated compact" in 13.8 kV, for remote

areas and network ducts or "cable rack" to load centers near the substation (plant).

For priority load areas, such as slurry pumps, thickener and tailings detoxification, an

emergency generator will be provided.



18.3.3 Fuel Supply System

There will be a gasoline and diesel station consisting of a number of tanks situated within a

bunded area with appropriate connections to dispensing pumps so control of consumption

can be affected. Diesel, and other fuel and lubricant supplies, will be contracted to a national

distributor. A preliminary cost estimate appropriate to the project’s requirements has been

received and is included in the capital cost estimates.

18.4 Communication Systems (Internal and External)

A voice and data communication system will meet the requirements of the project and

company, both internally and externally.

The communications center will provide services such as voice mail, calendar, automatic

transfers, automatic dialers, and facilities to support the operations and maintenance, such as

radio, intercom, fault diagnostics, and interconnection to the corporate network.

It is planned to install a radio communication system, using VHF frequency, with 100%

coverage of the process areas. Fourty units are proposed of the portable Motorola EP450

battery tyre with rapid charger and belt clip. This will provide a mobile communication system

for vehicles and field personnel during operations, maintenance and security.

18.5 Buildings - Maintenance Workshop, Office Buildings and Restaurant

18.5.1 Gatehouse

This is located on the side of the road entrance to the site and will compose a reception booth

where identity of any visitor will be checked before being given an electronic card which

enables the visitor to enter through the turnstile type gate. There will also be room where

visitors waiting can be shown videos of the general safety regulations applicable in the plant.

Under normal circumstances vehicles will not be allowed to enter the site and will have to be

parked outside the main gatehouse. Supply or maintenance vehicles which have to enter the

industrial area will be thoroughly checked before and after leaving. In the case of vehicles

carrying reagents the drivers will have to get weighed on the road weighbridge before and

after delivery with the weigh tickets handed in to the Gatehouse guards.

18.5.2 Main Office Buildings

All administrative activities including local financial management, human resources and

purchasing will be housed in the main office block. There will also be offices for the main

management posts such as General, Mine and Process managers as well as offices for the

environmental, information and security departments. There will also be an attended reception

office which will act as telephone exchange.



18.5.3 Central Restaurant

The central restaurant will be located within easy reach of the majority of employees. This will

be operated 24 hours a day so that all shift employees will also use the facility. The company

will supply a well equipped kitchen area including appropriate storage and freezer areas, but

the service of supplying the required food and its subsequent preparation and distribution will

be carried out by a contracted company. The size of the restaurant will be calculated for 120

people so as to cater for the larger demand during lunch-time during the day.

18.5.4 Nurses Clinic

This will be manned 24 hours per day by qualified nurses and have sufficient equipment for

emergency first aid and will also have an ambulance on permanent stand-by. The driver of

this vehicle will be a trained member of the operating staff on each shift.

18.5.5 Stores

The stores will attend the necessities of the plant with a large fenced outside area where

larger items such as crusher mantles, etc. will be located. The covered area will be of simple

construction using a metal roof with brick walls and ample shelving located along aisles.

18.5.6 Central Maintenance Workshop

This workshop will be a covered shop with walls. It will attend to maintenance and repairs of

all the mechanical equipment (both mine and processing plant) and have appropriate

equipment and tools such as welding, drill machines, a lathe as well as a small overhead

crane.

It is planned to provide a centralised compressed air facility in the workshop, with mobile

compressors for the mining area and additional compressors for the explosives factory.

All equipment will be provided with secure locking systems to ensure safety standards are

met.

Management and administration of the maintenance workshop will be located in offices within

the workshop.

18.5.7 Sanitary Waste and General Waste Disposal

Sanitary waste generated in all the administration and operational areas will be treated in

independent systems comprising of septic tanks and anaerobic filters.

The maximum of recycling will be carried out with cooperation with local recycling

installations. Organic waste will be shipped to a sanitary landfill. Naturally all packaging

associated with dangerous chemicals (cyanide) will be destroyed.



18.6 Explosives Magazine

The explosive storage facility was designed according to Brazilian regulations and contains

eight separated bunkers capable of storing a total of 1,800 t of ammonium nitrate which will

include a combination of ANFO (Ammonium Nitrate and fuel oil) and of emulsion explosives.

In addition, there is another bunker for the storage of dynamite, blasting caps and primers.

This area is located far enough from the mine (approx. 0.5 km) and from support facilities

(approx.1.6km) to meet the provisions of Brazilian regulation.

18.7 Provisional Facilities (Implementation Period)

An area of 120,000 m2, with a central maintenance workshop, will be allocated for the

construction site and temporary facilities of the contractors. Amarillo will require temporary

premises for supervision and monitoring of project implementation.

The accommodation will be located near the plant, and may be used with appropriate

modifications during the life of the mine.

In principle, to the extent that the permanent premises of the project are completed and

"delivered" to Amarillo, the relevant temporary facilities will be totally disabled, or demolished.



19 MARKET STUDIES AND CONTRACTS

19.1 Industry Trends and Pricing

The gold price in 2011, achieved a record spot trading value of more than US$1,800 per

ounce.

This increase in the price of gold appears to have been brought about by the uncertainty of

the strengths of the world currencies, the lack of economic recovery in Europe and U.S.A. that

was predicted to occur in 2010, the uncertainty in the mid-East where the majority of the oil

reserves lie, and the prediction of the rapid growth of inflation of newly developing countries

such as the BRIC countries. The expectation is that the price of gold will remain strong for the

near future because of the investment demand by ETFs, fabrication demand, purchasers of

gold coins, and banks.

The gold price that was utilized for the base case cash flow analysis is US$1,100 per ounce

for the life of the project, which approximates the three year trailing average for gold.

19.2 Sale Strategy

The Posse Gold Project will produce gold bars containing about 95% gold. These bars will be

refined to produce pure gold and the refined gold will be sold to banks or other financial

institutions either in Brazil or offshore on a spot price basis to capture the highest price.



20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

The most recent environmental studies undertaken for Amarillo Gold Corporation’s Mara

Rosa Project were undertaken by Neotropica Technologia Ambienttal Ltda. in 2010 and 2011

(Neotropica, 2010, 2011).

Information from these studies were used to assess the projects potential impact, if any, on

the environment along with identifying the necessary permit requirements for the project.

20.1 Physical Environment

The climate for the Mara Rosa project site was obtained from the Porangatu station, number

32346, located in the community of Porangatu, approximately 90 km from the study area.

Data was obtained for air temperature, relative humidity and precipitation.

Generally, the climate for the region is characterized by two distinct periods of the year: one

cool and dry, and the second hot and rainy. Average temperatures recorded in May and June

are about 24 °C and average temperatures are in August and September reach 28 °C.

Average rainfall as measured at Estrela de Norte (30 km to the north) is 1,679 mm (period

1971 to 2010). The relative humidity, for 2009, presents the highest values in April (84%) and

December (82%) and the lowest values are in July and August, reaching 45% and 40%

respectively.

20.2 Natural Environment

20.2.1 Terrestrial Environment

The natural environmental studies were undertaken to establish baseline conditions for flora,

mammalian, reptiles and amphibians and birds, hydrology and hydrogeology at the Mara

Rosa Project site that can be used later in identifying the possible impacts generated by the

Amarillo Gold Mining Company in the area that pertains to it, located in the municipality of

Mara Rosa, Goiás. This is an essential step towards the planning of future efforts that look to

increase conservation and environmental quality in the area surrounding Amarillo’s Project

site and in adjacent regions. The state of Goiás is located in the Cerrado Biome. The

cerrado, an immense tropical savannah, constitutes Brazil’s second largest plant formation.

This Biome is found predominantly in Brazil’s Central Plateau and covers 23% of this nation’s

territory. It should be noted that the Cerrado Biome is considered one of the 25 places on

Earth that have high biodiversity and are most threatened (Mittermeier et al., 1998; Myers et

al., 2000). According to Mittermeier et al. (1998), approximately 50% of all the terrestrial

biodiversity is found in these 25 areas; however, these 25 areas represent only 2% of the total

surface area of the planet. Recent estimates (Dias, 1992) show that approximately 50% of

the native vegetative cover has been destroyed, being that less than 5% of the total area of

the Cerrado Biome is protected in the form of conservation units. The high rate of occupation

of the Cerrado Biome, in an attempt to transform it into “the country’s great breadbasket”, is

the result of various socioeconomic causes.



Based on a review of satellite imagery, four distinct vegetative communities were identified as

present on the Mara Rosa Project site. Biological studies were undertaken in each of these

four communities.

Flora

The vegetation survey undertaken for the Mara Rosa Project site was to evaluate whether or

not there are plant formations that are of special interest and, if so, how well preserved they

are. The ecological functions and the environmental services that the vegetation performs, its

floristic composition, and whether or not it contains species that are common or protected by

law, were also evaluated in the study. The aim of this study was to provide a basis for the

constitution of a database that will complement and improve the information that is at the

mining company’s disposal so that it may plan a sustainable method of ore extraction. This

study is also intended to aid in the development of mitigation measures and/or the creation of

genuine areas of conservation for the natural fragments that lie in the region that will be

preserved or recovered.

The area that was sampled in this study is made up of a mosaic of vegetation that includes

strict sense cerrado vegetation, which dominates the landscape, and the cerradão variation,

or forested savannah, a forest typical of regions containing cerrado vegetation whose trees

can grow to be 15 m tall. Areas strictly representative of forests were also observed, such as

gallery forests, which, in the region, occur frequently and constitute a form of transitionary

vegetation between the forest canopy and the strict sense cerrado vegetation.

Based on the results of the vegetation surveys it was concluded that the reactivation of the

mine’s ore extraction area will not affect many natural areas, being that the place that will be

occupied by the Amarillo Gold Mining Company’s future installation has already been altered

by other human activities, such as pastureland and sansão do campo (Mimosa

caesalpineafolia) shrubs. The native plant formations of special interest are present on site

but are restricted to the riparian forests which follow the water courses in the area of

influence. It appears that future disturbances due to the extraction of ore will occur in this

vegetative community. This is already home to the greatest diversity of species whose

primary function is the conservation of soil and water resources, in addition to having other

protected vegetation species. In the case of the legal reserve area, the affected species

should be relocated to another location having similar vegetation on the same property if

possible and if not, within the same watershed. This process must be carried out in

accordance with procedures laid out in Ordinance No. 14/2001 and 15/2001 by SEMARH

(Secretariat of the Environment and Water Resources of the state of Gioás).

Mammals

The community of mammals in the Cerrado Biome is composed of 185 species, being that, of

these, 19 are common and 13 are officially considered by IBAMA (the Brazilian Institute of

Environment and Renewable Natural Resources) to be threatened by extinction (2003). The

mammals of the Cerrado present a low degree of endemism: 51% of the species found in this



biome are also found in the Amazon Biome; 38% in the Caatinga Biome; 49% in the Chaco

Biome, and 58% in the Mata Atlântica (Atlantic Forest Biome) (Alho et al., 1986).

During the course of the study, 16 total species of mammals were logged through methods of

direct recognition, the collection of indirect evidence, and by conducting interviews. All of the

mammal species that were recorded in the inventory are widely distributed geographically,

and most of them occupy the entire Cerrado Biome as well as sections of many other

Brazilian biomes. According to the List of Species of the Brazilian Fauna that are Threatened

with Extinction, published by the Ministério do Meio Ambiente/IBAMA (Ministry of the

Environment/IBAMA) (Brazil, 2003) and by the International Union for Conservation of Nature

(IUCN, 2010), two species that were mentioned in the interviews are considered vulnerable to

extinction: the Maned Wolf (Chrysocyon brachyurus) and the Giant Anteater (Myrmecophaga

tridactyla). This confirms the need for long-term studies that make use of various sampling

techniques in order to obtain quantitative data on biodiversity.

Reptiles and Amphibians

In Brazil there are 721 known naturally occurring reptile species (Brazilian Society of

Herpetology, 2010). Of these, only 20 are considered threatened (Rodrigues, 2005).

Considering exclusively the Cerrado Biome, the number of reptile species that are endemic to

it, and their quantities, are the following: 8 species of amphisbaenia (corresponding to 50% of

the total species of amphisbaenia), 12 species of lizard (26%), and 11 species of serpent

(COLLI et al., 2002; BASTOS, 2007). According to Colli et al., (2002), 3 amphibian species, 4

turtle species, 5 crocodilian species, 5 lizard species, and 6 serpent species that inhabit the

Cerrado are threatened with extinction.

Twenty-two species of reptiles and amphibians were recorded during the course of the

inventory on the Mara Rosa project site. In general, the diversity of the herpetofauna was

within what was expected relative to the degree of conservation of the area, the methods that

were adopted, and the meteorological conditions in which the study was conducted. All of the

species that were recorded can be found throughout the Cerrado Biome, and as none of them

appear on any lists of species that are threatened with extinction, none of them require close

attention. The data that were gathered for this inventory, although satisfactory, are too few

and lack the sophistication necessary for the making of inferences regarding population

estimation and the conservation status of the region.

Birds

Birds have the highest number of species that are described among the terrestrial vertebrates

registered for Brazil. In Brazil, 1,832 species have been catalogued (CBRO – Brazilian

Committee of Ornithological Registers – 2010). Silva & Bates (2002) have described 837

species within the Cerrado Biome; of these 30 are common to the Biome.

The bird surveys undertaken for the Mara Rosa project site resulted in the cataloguing of

seventy-nine bird species. This number is considered low, illustrating that further studies are



necessary to gain a better understanding of the local bird community. It should be noted that

the region of the future mining project is host to a considerable degree of human activity such

as large expanses of pastureland, areas where crop cultivation occurs (saffron, rice, etc.), and

the disturbances that were caused by the mining company that operated on the property

formerly and that according to Silva (1995), who conducted a major study of bird species

found in the Biome, the birds of the Cerrado, they are highly dependent on forest

environments. As such, the monitoring of the species of this Class is necessary in order to

quantify and qualify the population levels of this group so that rapid changes in the

environment, caused by human activity (e.g. mining projects), do not cause a great

disturbance in the avifauna communities. Long-term studies are fundamental in analyzing the

effects that are engendered by the installation of mining facilities, generating information and

answers concerning the integrity of the populations of these species over extended periods of

time.

20.2.2 Aquatic Environment

The aquatic environment study area included 3 sub-basins. These were: Upstream Basin for

the Ouro River; the sub-basin for Lambari Creek and the sub-basin for the Antas River.

Surface Water Quality

Surface water quality parameters were selected as being the most relevant for

characterization of natural water quality, since they are those which most greatly influence the

ecological standards for aquatic communities. Additionally, they are the most effective for

assessment of quality for direct and indirect human use. The analytical methods applied, and

the proposed legislative standards followed CONAMA N° 357/05, class II.

The surface water results for all points, including those in the pits are within the limits of class

ll of CONAMA Resolution n° 357/2005. The accumulated water in the pits is in general from

rainwater, beyond having a portion drawn from interstitial aquifers. Considering art. 34 and

Table X of CONAMA Resolution 357/2005 (that allows the direct or indirect discharge of any

potential polluting source in to water bodies as long as they meet defined requirements), the

water in the pits is satisfactory and can be discharged to the waters in the area of the project.

Phytoplankton and Zooplankton Communities

A qualitative study of the phytoplankton and zooplankton communities (indicators of water

quality) was carried out at 2 sampling points corresponding to pits 01 and 02.

The phytoplankton community is influenced by variations in temperature, pH, nutrient

concentration, hydrodynamics, as well as predatory action. An important characteristic of

phytoplankton is their rapid response to environmental change, as a result of their short life-

cycle, which makes them a very efficient indicator of water quality (Round, 1993; Reynolds et

al., 2002). The analysis of the phytoplankton community sampled in January 2011, in the

area of influence of the proposed project of Amarillo Gold demonstrated a relatively high

taxonomic complexity with an inventory of 34 taxons, distributed across 6 taxonomic groups.



Chlorophyceae was the group most highly represented, followed by Bacillariophyceae,

Zygnemaphyceae and Cyanobacteria.

Zooplankton also performs an important role in aquatic environments, forming a link in the

food chain between phytoplankton and other animals (Nogueira, 1996). To summarize, the

composition, richness and abundance results, obtained for the zooplankton, in the old mine

pits, in the area of influence of the mining project of Amarillo Gold, currently filled with

rainwater and spring water, presented a typical tropical freshwater state for high retained

water, with trophic conditions tending toward oligotrophic, taking into account the relatively

low community density. It must be noted that the higher diversity of testacean protozoans

suggests an influence coming from the littoral region on the structure of the zooplankton

community in the study area.

Benthic Macro-invertebrate Community

A qualitative-quantitative study of the benthic macro-invertebrate community (indicator of

water quality) was carried out at 2 sampling points corresponding to pits 01 and 02.

In regards to the benthic community, studies of the community structure have become

essential in evaluating environmental impacts of aquatic ecosystems. Changes in community

organization provide important information when the objective is bio monitoring of these

systems (Callisto, 2001). The benthic community was found to be extremely poor and the

taxa living in the pits are extremely generalist by nature, found in most aquatic environments.

As such, drainage of the pits would not cause problems of general diversity, as they are holes

having still waters and no connection to other water bodies, in addition to not having any

sensitive species, only resistant ones.

Sediment Quality

A sediment quality sampling program was carried out at 2 sampling points corresponding to

pits 01 and 02.

Heavy metal concentrations in sediments were measured at one sample location (AS-01).

Based on these sediment quality results, sediment at the bottom of the ponds should be

transferred to the reject dam to aid in reducing the possibility of contaminating local surface

water courses. Relevant legislation, CONAMA N° 344/04, allows the disposal of dredged

sediments, in water in Brazilian jurisdictions, without the necessity of additional

characterization studies as long as the concentration of pollutants is less than or equal to

Level 1, or for material where the concentration of metals, except mercury, cadmium, lead or

arsenic, between Levels 1 and 2.

Continuous monitoring is recommended for surface waters and aquatic communities at all

sampling points, with the exception of those related to the pits (AS-01 and AS-02) which will

likely be drained. Any expansion of the plant area should include an increase in the number

of surface water sampling points in order to take in the entire area of influence.



Water – Hydrology and Hydrogeology

The proposed area of the Mara Rosa Project mine and plant shows obvious signs of previous

mining. The last time mining was carried out in the area was in 1995. As a result the

environment shows significant changes from what would have been its previous natural state

including the presence of waste and tailing heaps only partially covered with vegetation.

Some deforestation is also visible although after cessation of mining activities, attempts were

made to ameliorate the situation and natural growth has made considerable headway in

hiding some of the worst aspects of the previous mining.

In addition the previous mining has left two small lakes where the surface mining had

penetrated up to 50 m below the original surface.

The updated Project Masterplan (or site layout, see section 18) shows the position of the main

features in the future mine, including the positions of the enlarged open pit, tailings dams,

waste heaps and industrial installations. Using this information a well defined plan for the

monitoring and sampling of ground and surface water has been developed to permit

understanding of the hydrogeology of the area (see H for full details). This data collection

along with its interpretation will allow Amarillo Gold to quantify flows, monitor water quality and

thus collaborate in satisfying any legal requirements.

It is noted that the area around the southern part of Mara Rosa town has markedly different

geology when compared with the northern part of the municipality around the region of the

Posse mine site, where crystalline rocks of poor water storage capacity predominate. The

result is that whilst the Mara Rosa town successfully draws its water supply from wells, very

little underground water is available around the mine site.

There is a lack of hydrological data with respect to the influence of seasonality of flows in the

streams and rivers in the area. The regulations state that only a certain quantity of water can

be extracted for industrial or agricultural uses. Information from flow monitoring points will be

required to establish if the mine’s processing plant requirements (40 m3/h) will be permitted

and that the State of Goias will therefore provide the appropriate license. In the absence of

this data, rough estimates have been made using values for potential aquifer production from

this type of rock and, considering the area which coincides with the watershed surrounding

the Rio do Ouro basin of about 95 km2. A probable value in excess of 626 m3/h has been

estimated, (after allowances for agricultural activities). This value is far and above what will be

required for the project and thus there is no danger that the project will not be able to function

by using local sources of water.

Previous analysis of ground and flowing water has not detected any values of contaminants

which are not within CONAMA Class II purity limits.

In relation to the environment, it is recommended that all existing wells used for monitoring

groundwater be blocked off since some of them are open to the atmosphere (allowing the

possibility of contamination from above). There are no construction details of these



monitoring wells and it is suspected that at least some have inadequate access to

groundwater.

It is possible that there also exists an environmental liability in connection with the tailings

deposited in the Baribras area (an area of 10 ha located within Amarillo Gold’s suspended

mining license, near the proposed open pit). These tailings were deposited when high

sulphide materials from the Zacarias mining area, situated about 10 km from the Mara Rosa

Project, were being processed in the pre-1995 processing plant. These tailings are

supposedly confined in a plastic lined area. However, it is thought that the high sulphide

content presents a potential risk of AMD (acid mine drainage). It is recommended that

samples are taken, and leaching and solubilisation tests be carried out to determine the

classification of the dump material and its potential for contamination (by acid generation).

Continued assistance will be required to provide correct diagnosis, assessments and plans to

mitigate the damage caused by the presence of the mine and plant especially with respect to

the areas bordering the streams and rivers (the so-called APP areas, areas of permanent

preservation). This information will be incorporated within an Environmental Impact Study

which will contain information from this and continued hydrological studies as well as fauna,

flora, socio-economic and other studies.

20.3 Social Environment

A preliminary socio-economic study was undertaken for the project to characterize the current

stage of human-influenced environmental factors in the direct and indirect areas of influence

of the project, which include: demographics; land use; production and economics; quality of

life, evaluated using indicators relative to education, health, human development index (HDI),

security; infrastructure related to communication, transportation, electricity, habitation and

sanitation, culture, recreation and tourism.

The methodology used in this study consisted of a survey of primary and secondary data from

the municipality of Mara Rosa and interviews with the directly affected owners of property

used for drilling and potentially extraction. The primary data was collected in the municipal

center and in the areas directly affected by the project. Secondary data was accessed, for the

most part, based on visits, via internet, to websites of official production bodies and/or

statistical information sources.

The municipality of Mara Rosa is located in the central area of the Tocantins River Basin. It is

part of the micro-region of Porangatu, the meso-region in the North of the state of Goiás. The

population of the municipality’s region was established in 1742, when Amaro Moreira Leite, as

leader of an expeditionary force, encountered a large quantity of gold during a river crossing,

later named the Rio do Ouro (River of Gold). The population count carried out by the

Brazilian Institute of Geography and Statistics (IBGE), for the 2010 Census, counted 10,659

inhabitants in Mara Rosa, with a population density of 6.32 inhabitants/km². Based on data

from 1991 to 2010, the population of the area has steadily decreased by about 50%.



Mara Rosa is situated 370 km from the capital and with an area of 1,703.95 km2, 69.84% of

which is urban area and 30.16% rural, its neighbours are the municipalities of Mutunópolis,

Estrela do Norte, Formoso, Campinorte, Nova Iguaçu, Amaralina, Santa Terezinha de Goiás,

Alto Horizonte, Campos Verdes and Uirapuru. The municipality is accessible by highways and

local municipal roads.

The economic base of Mara Rosa is rooted in agriculture and livestock. With few exceptions,

the producers of Mara Rosa have as a principal activity the cultivation of saffron, the base of

their economy. Cultivation of corn, beans, rice and fruits appear only as supplementary

income. The same happens with dairy cattle, chickens and pigs which are occasionally sold.

Some people also obtain income from non-farm-based sources. Immediate to the Mara Rosa

project site there are four owners that lease property to Amarillo Gold Corporation. Two of the

four property owners were interviewed, they are married and live with their families; they have

no employees contracted to assist with livestock and farming. According to the survey the

family grows crops for production for local commerce and consumption, as with beef cattle,

dairy cattle and pigs. It is significant to mention that the neighbouring properties to the future

project property are of small size, being between 5 ha and 69 ha.

The area is well serviced by police, fire, a number of health facilities and schools. The Mara

Rosa area is considered to have high human development when compared to other areas in

Brazil. Potable water and sanitary services are available for much of the Mara Rosa area but

not all residential homes have them. The area is also serviced by radio, TV and a newspaper.

If the exploration activities currently underway prove that the Mara Rosa project is viable, a

Social Communications Program will be required, focussed on establishing permanent social

communication channels with the community and local authorities, in order to reduce

uncertainty and improve the projects community image. The homeowners interviewed believe

that the intended commissioning of the project could contribute to socio-economic

development in the region, mainly in the generation of employment and income. The increase

in the number of vehicles and people in the municipality will require infrastructure in

installation and maintenance of roads, making the current roads and routes better from the

points of view of security, social accessibility and erosion control, taking into account the

ruggedness of terrain.

In the future, the operations phase of the mining project will require improvement in

communication and electrical distribution infrastructure. For this, the installation of electrical

substations and water treatment stations, among others, are expected, which will contribute to

the development and quality of life of employees and indirectly for the general municipal

population. These projects should be duly licensed, along with the processing plant.

In the case of commissioning/operation of a mineral extraction project there is a necessity for

environmental licensing, and for this a more detailed archaeological report/survey would be

required.



20.4 Waste and Tailings Disposal, Site Monitoring and Water Management

20.4.1 Mining Area

Open-pit mining, the method of extraction proposed, will potentially create zones of

geotechnical instability, where landslides and the crumbling of soils may occur. To avoid

and/or minimize the incidence of these events, the opening of the pits should be conducted in

observance of adequate technical criteria. A system of pumps is proposed to be constructed

at the pit bottom in order to keep the pit dry. This water will be pumped to the lake of

untreated water (WSF) for use in the processing plant.

20.4.2 Tailings Basins

The walls of the tailings basins will be constructed with the waste that is produced during the

initial mining stages and lined with compacted material. The pulp will settle, after which a

liquid surface will form. The humidity of the solids at the bottom of the basin will reach levels

of 25%. The walls of the basins will be raised in sequential steps during the life of the mine

and will always have a height just above that of the level of tailings in the basin. The water

will be pumped, from pumps attached to a barge floating on the surface of the tailings, back to

the process tank to be re-used. The level of the tailings basins will be controlled by both the

system of pumps and evaporation. The tailings basins, constructed in areas that are adjacent

to the mineral processing plant, will shoulder the burden of neutralizing the chemical

compounds that are used in processing of the ore. The floor of these dykes and basins

should be rendered impermeable in order to impede the movement of potential contaminants

to the environment.

20.4.3 Water Usage

Approximately 5,000 m3 of water will be consumed per day. The two sources that will supply

the majority of the water that will be consumed are:

the water that will be pumped from the pit, and

the water that is to return from the tailings dam.

Although water from wells and streams will also have to be used, this practice will be the

focus of a more detailed hydrological study. In the dry season, evaporation is a factor that will

clearly limit the amount of water that is supplied by the two main sources discussed above.

As such, the water that is consumed from wells and streams will vary greatly, but will be quite

limited during the rainy season.

The process water tank will consist of a stell tank with a capacity of 1,000 m3.

20.4.4 Site Monitoring & Water Management

The indicators that should be monitored on the Mara Rosa Project site regarding the objective

of maintaining low particulate matter and vehicle engine emissions is air quality monitoring

during the entire implementation phase of the project; with regard to the objective of



minimizing soil erosion and the removal of unconsolidated materials, water quality has been

chosen as the indicator that should be monitored during the entire implementation phase of

the project; and concerning the objective of restricting anthropic interventions outside the

service area, the indicator that has been selected is “Number of studies approved for

interventions in restricted areas”, such as those that lie close to the margins of watercourses

and those that are significant from an environmental-floristic viewpoint.

20.5 Permitting Requirements

In Brazil, mining generally is regulated by a series of sets of legislation, with the three levels of

government office having authorities over mining and environmental issues. At the federal

level, the organizations that set regulations and guidelines, as well as granting, monitoring

and enforcement of mining and environmental legislation for processing of mineral resources

are the following:

Ministry of the Environment – MMA: responsible for formulation and coordination of

environmental policy, as well as monitoring and supervision of implementation;

Ministry of Mines and Energy – MME: responsible for formulation and coordination of

minerals, electrical and petroleum/gas sector policies;

Secretary of Mines and Metallurgy – SMM/MME: responsible for formulation and

coordination of the implementation of minerals sector policy;

National Mineral Production Department – DNPM: responsible for the planning and support

of processing of mineral resources, preservation and study of paleontological heritage, also

supervising geological and mineral research, as well as granting, controlling and enforcement

in mining activities in all national territory, in accordance with the Mining Code;

Brazilian Geological Service – CPRM (Companhia de Pesquisa de Recursos Minerais):

responsible for generation and dissemination of basic geological and hydrological knowledge,

in addition to making available information and knowledge about the physical environment for

territorial management purposes;

National Water Agency – ANA: Responsible for the implementation of the National Water

Resources Policy, its main authority being implementation and management of the country’s

water resources. Responsible as well for the granting of surface and ground water rights,

including those used in mining.

National Environmental Counsel – CONAMA: responsible for formulation of environmental

policy, resolutions with regulatory power, under law, whenever the legislative authority has not

approved specific laws;



National Water Resources Counsel – CNRH: responsible for formulation of water resource

policy; promote the coordination of water resources planning; establish general criteria for

granting of water use rights and charges for their use.

Brazilian Institute for Renewable Natural Environmental Resources – IBAMA:

responsible, at the federal level, for environmental licensing and enforcement;

Cave Study Centre – CECAV (IBAMA): responsible for caves and heritage.

According to the “Guia do Minerador – 2000” constitutional legislation, which directs

environmental policies and laws relative to the activities of mining, is basically consolidated in

the following legal documents, resolutions and ordinances:

Federal Laws:

Law nº 6938, of 31 August 1981 and its revisions (Law nº 7804, 18 July 1989, and

8028, 12 April 1990) – provides National Environmental Policy, its purposes, means

and application;

Law nº 9537, 11 December 1997 – provides direction water transportation in waters

under national jurisdiction and designates the Maritime Authority to establish standards

for construction, dredging, research, and mining, on, in and on the margins of waters

under Brazilian jurisdiction.

Federal Decrees:

Decree nº 97632 of 10 April 1989 – provides for recovery of areas degraded by mining;

Decree nº 99274 of 6 June 1990 – Regulates Law nº 6938, 31 August 1981.

Resolutions of National Environmental Counsel - CONAMA

CONAMA Resolution nº 01, 23 January 1986 – Establish basic criteria and general

directives for Environmental Impact Studies (EIA) and Environmental Impact Reports

(RIMA);

CONAMA Resolution nº 009, 6 December 1990 – Provides specific standards for

obtaining an environmental license for mineral extraction, except as related to

construction.

CONAMA Resolution nº 010, 6 December 1990 – Establishes specific criteria for the

extraction of mineral substances for immediate uses in construction.

CONAMA Resolution nº 2, 18 April 1996 – Provides details relating to compensation

for environmental damage, caused by projects and relevant environmental impacts;



CONAMA Resolution nº 237, 19 December 1997 – Provides details about procedures

and criteria to be used in environmental licensing.

CONAMA Resolution nº 303, 20 March 2002 – Provides parameters, definitions and

limits for Environmental Preserves.

Currently, Amarillo Gold Corporation has engaged a local Brazilian Environmental company to

undertake all environmental permitting requirements for the project. It is Coffey’s

understanding that currently there are no outstanding permits for the project.

20.6 Mine Closure

EIA and RIMA reports are being prepared by a specialist company in order to obtain the

appropriate environmental licenses. In Brazil, after the acceptance of these reports an LP,

(Preliminary License) is issued by the environmental authority. This is usually granted with

some conditions. A further report is then submitted which must comply with the conditions,

upon which an LI (Installation License) is granted. After construction is completed a final LO

(Operating License) is granted if the construction has been in accord with previous

agreements and the company is then free to operate (always obeying all the conditions

stipulated in the reports and the licences).

A fundamental part of this licensing process is a Mine Closure Plan. This is in the process of

being formulated but there follows a general view of what is intended for the closure of the

mine.

At the end of the mine life, the facilities will be completely dismantled and where relevant

equipments will be sold or removed as scrap. All concrete foundations will be buried and the

building sites graded and re-vegetated with indigenous species. Topsoil will be applied in

areas where it is considered necessary to enable successful re-vegetation.

During the installation of the mine (formation of tailings dams, waste piles, etc.) topsoil will be

removed, fertilized and vegetated with native species and kept in clearly marked areas for use

during mine closure.

A preliminary assessment of the work and material quantities involved in TSF closure has

been completed together with the TSF design work. Further study to fully cost the closure

operations will be undertaken as part of feasibility studies.

The level of the water reservoir will be monitored during the final phases of the life of the mine

so as to have ample capacity to receive the supernatant from the tailings dam (which can be

done in phases if necessary). The water within the water reservoir will be monitored until a

total cyanide content of less than 0.05 ppm is obtained. This may take some time (possibly

two years) and may involve the application of a detoxification process to help natural

degradation by ultra-violet rays from the sun. When the level of toxicity of the water reservoir

is regarded as safe then the water can be used for agricultural purposes.



The waste rock stockpiles will be contoured, covered with soil, and planted with native

species to prevent erosion. Water emanating from the waste heaps will be directed using

ditches to sediment ponds to allow fine sediments to settle prior to direct discharge into the

environment.

The mine pit will be allowed to flood, (the current small pits left from previous operations are

also currently filled with water and contain fish and other aquatic species) and surrounded by

fences to prevent animals entering the area. The water quality in the pits will be monitored for

a period of at least two years.

The company will maintain a constant presence for at least 5 years, or until the environmental

authorities are satisfied that a stable situation has been achieved. A total value of US$8.3 M

has been provisionally earmarked for the shut-down phase but intended procedures must of

course be acceptable to the local environmental authorities. The estimated salvage value of

mine equipment is currently shown to exceed this closure cost estimate (Section 22).

A formal closure plan and cost estimate will be developed as part of the feasibility study work

plan for the Project.



21 CAPITAL AND OPERATING COSTS

21.1 Mining Capital Cost

Capital cost estimates were developed as follows:

Estimation of fleet equipment for the mine;

Timing of acquisition of mine equipment, including the initial capital cost;

Calculation of capital cost of mining equipment, based on unit prices provided by

suppliers; and

Review of capital costs of mining facilities and services estimated by Amarillo and Onix..

21.1.1 Equipment Cost

Manufacturers and suppliers provided the price of each piece of equipment.

The cost data that were provided include charges for preparing the equipment for immediate

use on site, such as shipping, insurance, tax and assembly costs. Coffey Mining did not

obtain independent price quotes for mine project equipment.

Prices were obtained from various suppliers through price quotes. Coffey Mining prepared an

initial list of equipment, based on the available price quotes. Coffey Mining considers this

equipment list adequate for the requirements of this estimate and believes that costs can be

improved upon final negotiations with the equipment suppliers.

21.1.2 Capital Cost for Mine Equipment

The initial capital cost and the additional investments required during years two and three

were calculated using the prices and technical information that were provided by suppliers, in

addition to the timetable for equipment acquisition for the mine. These data are shown in

Tables 21.1.2_1 and 21.1.2_2 below

The initial investment for mine equipment totals US$24.62 M. The additional capital that will

need to be invested in equipment during the life of the mine is estimated at US$8.62 M.

Total capital cost for mining equipment during the life of the mine is estimated at US$32.60 M.



Table 21.1.2_1

Mara Rosa Project

Mine Equipment

Equipment Life Cycle

capacity unit Equip. (qty)

Total Unit Cost Phase

Total h year Inv Reinv. US$ Inv Reinv.

Main Equipment

Excavator 6 m3 45,000 8 5.8 m3 2 0 2 1,290,000 2,580,000 - 2,580,000

Excavator 10 m3 45,000 8 10.5 m3 2 0 2 2,660,000 5,320,000 - 5,320,000

45t haul truck 45,000 8 40.0 t 2 1 3 1,220,800 831,579 415,789 1,247,368

100t haul truck 45,000 8 95.0 t 6 7 13 1,138,700 6,832,200 7,970,900 14,803,100

Wheel loader 45,000 8 4.5 m3 2 0 2 1,019,200 2,038,400 - 2,038,400

Pneumatic rock drill 25,000 8 4.0 in 2 0 2 426,489 852,979 - 852,979

Hydraulic rock drill 25,000 8 5.0 in 3 0 3 260,395 781,184 - 781,184

SubTotal 19 8 27 19,236,342 8,386,689 27,623,031

Ancillary Equipment

Track-type Tractor 45,000 8 410 Hp 2 0 2 1,517,600 3,035,200 - 3,035,200

Wheel Tractor-Scraper 45,000 8 285 Hp 1 0 1 739,200 739,200 - 739,200

Water truck 6x4 45,000 8 20,000 L 1 0 1 152,632 152,632 - 152,632

Service Truck 45,000 8 6 t 2 0 2 61,150 122,300 - 122,300

Truck Loading Crane* 45,000 8 3 t 1 0 1 - - - -

Flat-bed Truck – 3 axles 45,000 8 40 t 1 0 1 63,291 63,291 - 63,291

LTM 1050 Crane 45,000 8 50 t 1 0 1 399,418 399,418 - 399,418

Mining Twin Cab 4 x 4 pickup truck 22,000 4 - - 4 4 8 43,107 172,429 172,429 344,859

Compact vehicle (VW Gol, Fiat Uno, etc.) 22,000 4 - - 4 4 8 15,166 60,663 60,663 121,326

Pumps 1 3 4 634,667 1,125,488 1,760,155

SubTotal 18 11 29 5,379,800 1,358,580 6,738,381

Total - - - - 37 19 56 - 24,616,142 9,745,269 34,361,411



Table 21.1.2_2 Mara Rosa Project

Investment Schedule (all costs in US$)

Equipment

Year

0 1 2 3 4 5 6 7 Total

units cost units cost units cost units cost units cost units cost units cost units cost units cost

Main Equipment

Excavator 6 m3 2 2,580,000 - - - - - - - 2 2,580,000

Excavator 10 m3 2 5,320,000 - - - - - - - 2 5,320,000

Haul truck 45 t 2 831,579 - - 1 415,789 - - - - 3 831,579

Haul truck 100 t 6 6,832,200 3 3,416,100 1 1,138,700 3 3,416,100 - - - - 13 14,803,100

Wheel loader 2 2,038,400 - - - - - - - 2 2,038,400

Pneumatic rock drill 2 852,979 - - - - - - - 2 852,979

Hydraulic rock drill 3 781,184 - - - - - - - 3 781,184

Ancillary Equipment

Caterpillar D9T Track-type Tractor 2 3,035,200 - - - - - - - 2 3,035,200

Caterpillar 160M Wheel Tractor-Scraper 1 739,200 - - - - - - - 1 739,200

Water truck 6x4 1 152,632 - - - - - - - 1 152,632

Service Truck 2 122,300 - - - - - - - 2 122,300

Truck Loading Crane* 1 - - - - - - - - 1

Flat-bed Truck – 3 axles 1 63,291 - - - - - - - 1 63,291

LTM 1050 Crane 1 399,418 - - - - - - - 1 399,418

L200 Mining Cab pick up truck Twin Cab 4 172,429 - - 4 172,429 - - - - 8 172,429

Compact vehicle (VW Gol, Fiat Uno, etc.) 4 60,663 - - 4 60,663 - - - - 8 60,663

Pumps 1 634,667 1 219,022 2 906,466 - 1,760,155

TOTAL 37 24,616,412 3 3,416,100 2 1,357,722 14 4,971,447 0 0 0 0 0 0 0 0 56 34,361,411



21.1.3 Pre-Production Services Cost

Pre-production cost estimates are included in Table 21.1.3_1. They will be undertaken by

Amarillo.

21.1.4 Mine Services and Installations

The capital cost data for services and installations were obtained by Coffey Mining, which

verified the consistency and the quantity of these amenities and the main costs that compose

these items.

Coffey Mining estimated the amount of earth that should be moved during pre-production and

was responsible for the tailings basin and water reservoir project designs.

The initial capital costs for mine services and installations are summarized in Table 21.1.3_1.

The initial capital cost investment in services and installations for the mine is estimated at

US$38.81 M, while the total investment for the life of the mine is estimated at approximately

US$48.55 M.



Table 21.1.3_1

Mara Rosa Project

Disbursement Schedule – Initial Capital Cost and Reinvestment

Item

Year

0 1 2 3 4 5 6 7 Total

(US$)

Pre-Stripping (2) 14,191,382 14,191,382

Main Equipment (1) 19,236,342 3,416,100 1,138,700 3,831,889 27,623,031

Drainage and pumping system (1) 634,670 219,022 906,466 1,760,162

Ancillary Equipment (2) 4,745,130 233,090 4,978,220

Total 38,807,524 3,416,100 1,357,722 4,971,447 48,552,795

Sources:

(1) Supplier Proposals

(2) Coffey Estimate



21.2 Plant Capital Cost

21.2.1 Civil

With reference to construction such as administrative buildings/workshops/laboratory and

similar constructions an average price per square metre currently practiced in the city of Belo

Horizonte was taken. For other civil constructions (bases of tanks mills, etc.) the quantity of

earthworks was calculated and a quantity of reinforced concrete also computed using unit

values currently in use in Belo Horizonte. A value of US$886/m3 of reinforced concrete was

used.

21.2.2 Mechanical Equipment

Equipment prices are derived from actual quotations in response to specifications sent out to

traditional suppliers. All major equipments were priced in this manner and actual quotations

can be found in the Onix report supplied in the appendix. It should be noted that the taxes

such as ICMS (VAT), etc. were included but the ICMS contribution to Goiás state (this tax is

divided between producer and consumer state) was not included as this can be negotiated as

exempt for new plants (orientation from a tax specialist).

21.2.3 Platework

To obtain the cost of this item, weights of the appropriate chutes, hoppers and tanks were

calculated from the basic arrangement drawings (with appropriate allowances made for the

addition of any rubber or abrasion resistant surfaces) and a total cost calculated using a unit

cost/kg. This unit cost was equivalent to R$15/kg or US$ 7.9/kg (plain steel) and includes

fabrication, painting, linings (when applicable) and installation.

21.2.4 Metallic structures

Again the approximate weight was calculated using the lay-out drawings and appropriate

costs computed by multiplying the weight by the cost of the particular section. Again

fabrication, painting and erection costs are included.

21.2.5 Piping

The cost of the piping required in the plant has simply been taken as 20% of total mechanical

equipment costs.

21.2.6 Electrical/Instrumentation Equipment

All the major hardware such as MCC’s, transformers, vfd’s etc. have firm costs quoted by

electrical suppliers. This includes values for automation of the plant as can be seen in the

ONIX report in the Appendix. An additional 25% was added to the sum of all this material to

allow for the purchase of cable trays, cabling and illumination, etc.



21.2.7 Cost of Installation (electrical and mechanical)

A value of 20% of the cost of all the equipment (automation, electrical , including cables, etc

plus cost of piping and mechanical equipment) was taken to allow for the cost of installation.

21.2.8 Cost of the Water Supply

This was regarded as a separate item and all the equipment costs, the HDPE piping and

welding and general installation costs were included in the final stated amount.

21.2.9 Energy Supply

The cost of the switching station at Porangatu as well as the power line up to the plant was

regarded as a separate item. An actual estimate was received from a company who is

certified with CELG to carry out the entire installation.

21.2.10 Construction Management

The value of this item was taken to be 12.5% of the value of the mechanical equipment.

21.2.11 Miscellaneous Items

In other cases factors were used. A value of 5% of the mechanical equipment total was used

to represent the costs of temporary installations required during the construction phase and

include additional values for provisional transport and food requirements during construction

as well as extras for uniforms, protective clothing etc. Estimates of the cost of transport and

food during construction were based on values sourced in the general area. Initial spare parts

were also taken to be 5% of the cost of the mechanical equipments. It should be noted that

the full salaries of plant and administrative personnel (and associated costs) were added to

the operating costs for 2013 and this totals more than US$10 M. This corresponds to the cost

of maintaining the company’s presence during construction as well as training of operators.

21.2.12 Main Equipment Costs

A list of the major plant equipment with associated costs in US dollars is presented in Tables

21.2.12_1, 21.2.12_2 and 21.2.12_3.

It should be noted that for imported items, a value of 8.8% ICMS, 9.25% (Cofins and PIS) and

a 40% import duty has been assumed, whilst specifically for the ball mills, an import duty of

22% followed by ICMS of 6% and PIS/Cofins of 9.25% has been assumed as indicated by

information from suppliers.



Table 21.2.12_1

Mara Rosa Project

Plant Major Crushing Equipment List and Unit Costs

Description Fabricator Value US$

Hydraulic Rock Breaker RHT 50,794

Vibrating Feeder/Grizzly Bercam 76,316

Jaw Crusher, Telsmith HD 1400 Bercam 460,536

Crusher Discharge Conveyor Bercam 20,430

Primary Screen Conveyor Bercam 53,956

Weightometer Ramsey 31,588

Electromagnet for Scrap removal Inbras 38,854

Metal detector Inbras 17,091

Vibrating Feeder Bercam 76,316


Conveyor under first stockpile Bercam 25,904

Primary Screen feed Conveyor Bercam 73,378

Primary Screen, 1.8 x 4.9m VG 6x16DD Bercam 47,378

Primary Screen and Cone Discharge Bercam 27,115

Secondary Cone Crusher, Telsmith 52SBS Bercam 605,272

Tertiary Cone Crusher, Telsmith 52SBS Bercam 605,272

Secondary Screen Feed Conveyor Bercam 524,799

Secondary Screen Oversize (to second stockpile) Bercam 46,746

Secondary Screen, 2.44 x 6.1m SM 8 x 20 DD Bercam 94,746




Final product Conveyor Bercam 141,807


Crusher Area Sump pump FLSmidth 16,113

Belt Feed Conveyors, (A, B, C and D) Tecnometal 127,507

Lime Silo and Dosing System Ducon Powder 127,851

Emergency Feed Conveyor, using loaders Simplex 50,993

Total 3,519,749 Note: For items imported a value of 8.8% ICMS, 9.25% (Finsocial and PIS) and 40% import duty has been assumed.

For the ball mills a specific Import Duty of 22% followed by ICMS of 6% and PIS/Confins of 9.25% has been

assumed.

There have been adjustments to the stated cost of imported items to allow for variation in exchange rate



Table 21.2.12_2

Mara Rosa Project

Plant Major Milling Equipment List and Unit Costs


Ball Mill Feed conveyor Simplex 201,131


Lime Area Sump pump FLSmidth 15,902

Primary Ball Mill Outukumpu 6,559,668

Secondary Ball Mill Outukumpu 6,559,668

Cyclone Feed Pump FLSmidth 203,481

Mill Area Sump pump FLSmidth 35,691

Cyclone Nest FL Smidth 147,378

Trash Screen Delkor 286,909



assumed.

There have been adjustments to the stated cost of imported items to allow for variation in exchange rate



Table 21.2.12_3

Mara Rosa Project

Plant Major Gravity, CIP and Elution Sections Equipment List and Unit Costs


Gravity Feed Pump FLSmidth 23,902

Gravity Feed Preparation Screen Minspec 95,857

Knelson Gravity Centrifuge Knelson 271,043

ILR 100 Batch Gekko 396,648

Gravity Sump pump FLSmidth 15,902

Pre-Oxidation Feed Pumps FLSmidth 73,798

Primary and secondary feed Samplers Outotec 13,669

Pre-Oxidation and Leach Agitators CDC 1,614,324

Pre-Oxidation Recirculation Pump FLSmidth 153,166

Pre-Oxidation Sump Pump FLSmidth 17,849

PA Oxygen System Oxair 1,162,294

Interstage Carbon Transfer Pumps Weir 152,804

Interstage Carbon Screens, plus spare Kemix 1,167,418

Loaded Carbon Screen Minspec 36,855

Carbon Sizing Screen Minspec 36,855

Carbon Safety Screen Minspec 159,651

Leach area Sump pump FLSmidth 17,849

Thickener (plus concrete tank) VLC 1,105,263

Thickener Underflow pumps FLSmidth 59,271

Agitators for Detox tanks CDC 161,903

Detox Area Sump Pump FL Smidth 17,849

Tailings Pumps FLSmidth 98,639

Tailings Sampler Outotec 30,755

Elution and Carbon Reactivation/Gold room COMO 4,395,222

Tailings Reclaim pumps KSB 13,273

Process Water umps KSB 19,080

Raw Water Pumps KSB 9,655

Potable water pumps KSB 3,961

Mine Dewatering pump, diesel KSB 11,318

Plant Compressors Sullair 86,821

Reagent Dosing Pumps Netzsch 37,905

Reagent Dosing Pumps Graco 4,135

Copper Sulphate Storage and Mix tank Fibrav 4,680

Sodium Metabisulphite Storage and Mix tank Fibrav 8,831

Hydrochloric acid Stock tank Fibrav 15,745

Laboratory Equipment Essa 224,081

Atomic Absorbtion spectrometer Perkin Elmer 47,638

Water Treatment Plant (ETA) Veolia 14,169

Emergency generator Estimate 207,059



assumed.

There have been adjustments to the stated cost of imported items to allow for variation in exchange rate.



Table 21.2.12_4 summarises the total plant capital costs, including all related construction

costs and the tailings storage facility.

Table 21.2.12_4

Mara Rosa Project

Plant Capital Costs

Year 0 Year 1 Year 2 Year 3 Year 4 Year 5

Crushing 3,519,749

Milling 14,041,416

Gravity, CIP and Elution Sections

11,987,136

Platework 6,389,329

Metallic Structures 4,932,324

Piping 5,909,660

Electrical/Automation 12,575,845

Fire Prevention System 96,225

Electrical and Mechanical Installation

9,606,762

Support Buildings including Laboratory

2,249,368

Plant Support Buildings 965,526

Excavations and Civil Foundations

6,253,272

Initial Stock of Spares 1,477,415

Light Vehicles and Maintenance Vehicles

747,368

Provisional Installations (and transport, etc) during Construction

2,498,467

Water Reservoir (Proposal from Constructor Fagundes)

1,034,667

Tailings Dam (Proposal from Constructor Fagundes)

7,114,203 3,178,842 3,178,842 3,178,842

Total Process Plant Capital Cost

91,398,734 3,178,842 3,178,842 3,178,842



21.3 Infrastructure Capital Cost

Table 21.3_1 summarises the total infrastructure capital costs not included in the items listed

for the plant.

Table 21.3_1

Mara Rosa Project

Infrastructure Capital Costs

Item Source US$

Supply and Water Distribution (1) ONIX 2,431,778

Electric Power line from Porangatu (64km) ONIX 7,894,737

Fuel Supply System Petrobrás 178,700

Communication Systems (Internal & External) Coffey 66,632

Mine Infrastructure, including heavy vehicle workshop Coffey 2,568,420

Maintenance workshop, restaurant, etc. ONIX 502,263

Explosives magazine Coffey 236,840

Infra-Structure 13,879,370

21.4 Indirect Costs

Indirect Costs comprise allocations for studies and construction management as well as

certain miscellaneous costs. These have been factored from ONIX and Coffey Mining

experience and/or estimated by Amarillo from work proposals.

21.4.1 Studies and Construction Management

Table 21.4.1_1 lists the items under studies and construction management. Freight was

costed as 3% of the total value of the mechanical and electrical equipments plus the

estimated costs of the platework, metallic structures, water distribution and piping. The value

of specialized technical assistance was taken to be 1.25% of the total mechanical and

electrical/automation costs. The first fill of reagents was calculated.

Table 21.4.1_1

Mara Rosa Project

Studies and Construction Management Costs

Item US$

Basic and Detailed Engineering 2,414,368

Definitive Feasibility Study and Confirmatory Testwork 1,631,579

Environment and Social Communications Programmes 473,684

Construction Management 3,693,538

Consultancy and Technical Assistance 507,852

First Fill of Reagents and Lubricants 3,524,764

Freight 1,783,551

Total Studies And Construction Management 14,037,336



21.4.2 Miscellaneous

Table 21.4.2_1 lists the miscellaneous items. Various values such as the cost of training of

the labour force, travel and owners costs, and geotechnical drilling (US$700,000) were

estimated.

Table 21.4.2_1

Mara Rosa Project

Miscellaneous CostsItem Source US$Training ONIX 184,210 Travel ONIX 210,526 Recruitment ONIX 105,263 Owner's Costs ONIX 2,200,000 Geotechnical Drilling Amarillogold 700,000 Miscellaneous * 3,399,999

*Values presented by ONIX refer to investments in the process plant during operations.

21.4.3 Sundry Items

Sundry items comprise:

A rebate of US$2,631,579 in each of Years 2, 3 and 4 from the power utility CELG for

financing the power line connection (itemised under infrastructure capital);

Insurance of US$421,053 estimated to be 1% of the cost of the mechanical and

electrical/automation equipment.;

Contingency of 10% on the plant capital costs, estimated at US$16,194,402; and

Initial Working Capital of US$5,475,218, estimated from two months operating expenses

in Year 3.

21.5 Sustaining Capital Costs

21.5.1 Mining Equipment

A limited amount of sustaining capital is reported for Years 1, 2 and 3 to finance the

expansion of the main mining fleet. Some replacement capital for mine ancillary equipment is

allocated for Year 3 and additional pumping equipment will be required for Years 2, 3 and 5.


The construction of the tailings storage facility is four phase process, with the initial

construction included as initial capital in Year 0. Phases two, three and four are scheduled for

Years 2, 4 and 5 respectively, and are reported as sustaining capital.

21.5.3 Closure cost

Mine closure has been discussed in Section 16.2.5 and 20.6. A formal closure plan and cost

estimate will be developed as part of the feasibility study work plan for the Project. For the



purpose of this study an estimate of US$8.5 M has been prepared based on the material

quantities presented in Section 16.2.5. This amount is significantly less than the estimated

salvage value of equipment discussed in Section 22.1.5. both closure and salvage costs

have been expenses for the financial analysis.

21.6 Schedule of Capital Costs

A capital expenditures schedule has been developed from the project execution plan. Mine

pre-stripping and indirect costs come from the PFS mine schedule and have been allotted as

Project Capital.

Plant and Infrastructure direct and indirect costs will be expended in Year 0 (2013). Site

preparation costs will be incurred in 2013. Although site work may begin upon approval of the

Project EIA in late 2012, this work will likely not be expensed until 2013. Owner’s costs have

been estimated by Amarillo based on their current cost structure and the anticipated Project

requirements.

Mine closure costs have been expensed in Year 8.

The capital cost schedule is provided in Table 21.6_1.



Table 21.6_1

Mara Rosa Project Capital Cost Schedule

Cost Category Year 0 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Total Direct US$ x 1,000 US$ x 1,000 US$ x 1,000 US$ x 1,000 US$ x 1,000 US$ x 1,000 US$ x 1,000 US$ x 1,000 US$ x 1,000 Mine Pre stripping 14,191 - - - - - - - 14,191 Main equipment 19,236 3,416 1,139 3,832 - - - - 27,623 Auxiliary equipment 5,380 - 219 1,140 - - - - 6,738 Total 38,808 3,416 1,358 4,971 - - - - 48,553 Plant Mill 14,041 - - - - - - - 14,041 Process equipment 15,507 - - - - - - - 15,507

Boiler 6,389 - - - - - - - 6,389 Steel structures 4,932 - - - - - - - 4,932

Piping 5,910 - - - - - - - 5,910 Electrical/automation 12,576 - - - - - - - 12,576 Fire fighting system 96 - - - - - - - 96 Assembly 9,607 - - - - - - - 9,607 Administrative buildings 2,249 - - - - - - - 2,249 Architectural support 966 - - - - - - - 966 Excavation and civil works 6,253 - - - - - - - 6,253 Spare parts 1,477 - - - - - - - 1,477 Light vehicles 747 - - - - - - - 747 Provisional facilities 2,498 - - - - - - - 2,498 Dam water 1,035 - - - - - - - 1,035 Tailings dam 7,114 - 3,179 - 3,179 3,179 - - 16,651 Total 91,398 - 3,179 - 3,179 3,179 - - 100,935 Infrastructure Water supply systems 2,432 - - - - - - - 2,432 Lt-Bay in Porangatu 7,895 - - - - - - - 7,895 Fuel/lubricant systems 179 - - - - - - - 179 Communication systems 67 - - - - - - - 67 Mine support buildings 2,568 - - - - - - - 2,568 Explosives magazine 237 237 Misc infrastructure 502 - - - - - - - 502 Total 13,880 - - - - - - - 13,879 Total Direct 144,086 3,416 4,537 4,971 3,179 3,179 - - 163,367 Indirect Studies and management 14,037 - - - - - - - 14,037 Miscellaneous 3,400 - - - - - - - 3,400 Rebate from CLEG - (2,632) (2,632) (2,632) - - - - (7,895) Insurance 421 - - - - - - - 421 Process contingency 16,194 - - - - - - - 16,194 Initial working capital 5,475 - - - - - - - 5,475 Total Indirect 39,528 (2,632) (2,632) (2,632) - - - - 31,633 Total 183,614 785 1,905 2,340 3,179 3,179 - - 195,001

Rounding has been applied



21.7 Mining Operating Cost

21.7.1 Overall Aspects

The operating cost estimate is based on the extraction of approximately 7,100 tonnes of ore

per day, and the removal of waste rock in accordance with what was presented in the

production program (see Section 16) for a period of 360 days per year.

The operating costs were calculated annually for each period during the life of the project.

These cost estimates were based on the amount of operating hours of the equipment, the unit

costs applied to the different types of equipment, personnel requirements and the Brazilian

unit costs for raw materials and consumables, services and labour. These costs do not

include taxes and contingencies. The following exchange rate was used: R$1.90 to US$1.00.

The costs that are directly attributable to the mine area are related to, first, the mining

operations that either feed the primary crusher or the stockpile, and, second, the removal and

hauling of waste rock to the waste rock piles (in conformity with the waste rock disposal

schedule). All costs associated with operating and maintaining the mine are included under

the assumption that Amarillo will be the mine operator. At this stage, only the blasting

operations are assumed to be conducted by a blasting contractor.

The following items were included in the overall administrative costs and thus excluded from

the mining cost estimates:

Light vehicle expenditures;

Office and camp cleaning;

Business travel;

Maintenance of the main road;

Insurance and assets;

Safety inspections;

IT and communications;

HR;

Sampling and analysis;

Outside consulting services;

Mineral exploration expenditures.

21.7.2 Basic Consumption and Cost

The fuel and lubricant costs correspond to prices that were quoted by Petrobrás, a major

Brazilian supplier of petroleum products. Quoted prices were US$1.02/L and US$3.79/L

respectively. The consumption rates for each piece of mobile equipment are based on



information that was supplied by their manufacturers. These data are presented in Table

21.7.2_1.

Table 21.7.2_1

Mara Rosa Project

Main Equipment Hourly Fuel and Lubricant Consumption

Equipment Capacity Unit Fuel Lubricant

Rock Drill Rig – Ore 4 in 60 0.3

Rock Drill Rig – Waste 5 in 35 0.3

Wheel Loader 5 m3 28 0.3

Hydraulic Excavator – Ore 5.8 m3 62 0.5

Hydraulic Excavator – Waste 10.5 m3 147 0.7

Truck – Ore 45 t 25 0.3

Off-highway Truck – Waste 95 t 87 0.9

Bulldozer 460 Hp 55 0.4

Motor grader 200 Hp 30 0.3

Water Truck 20,000 L 18 0.2

Service Truck 6 t 12 0.2

Truck Loading Crane 3 t 18 0.2

Flat-bed Truck 40 t 20 0.2

Crane 50 t 25 0.2

An estimate of the consumption of drill bits, drill collars and stabilizers is presented in Table

21.7.2_2.

Table 21.7.2_2

Mara Rosa Project

Drilling Equipment – Costs and Life Cycle

Drilling Equipment Parts Cost

Rock Drill Rig- Waste Rock Drill Rig - Ore

Life Cycle

(US$) (m)

Bit 1,685.82 1,000 1,000

Collar 2,223.93 4,500 -

Drill Rod 1,466.68 1,500 -

Hammer 8,419.15 - 4,000

Crossover 860.22 - 8,000

Drill Tubes 1,742.53 - 8,000

Adapter 1,290.33 - 8,000

Rotation Unit 1,290.33 - 8,000



These costs are based on price quotes that were provided by ATLAS COPCO.

Tire wear estimates are based on data from gold ore mines of similar layout.

The wear parts that are associated with excavators, front loaders, tractors and wheel tractor-

scraper are summarized in Tables 21.7.2_3, 21.7.2_4 and 21.7.2_5.

Table 21.7.2_3

Mara Rosa Project

Excavator Wear Parts - Life Cycle and Costs

Part

Hydraulic Excavator - Ore Hydraulic Excavator - Waste

Life Cycle Cost Life Cycle Cost

(h) (US$) (h) (US$)

Bucket 5,000 18,500 20,000 120,600

Teeth 200 2,600 5,000 64,000

Table 21.7.2_4

Mara Rosa Project

Front Loader Wear Parts - Life Cycle and Costs

Part

Wheel Loader

Life Cycle Cost

(h) (US$)

Bucket 10,000 10,600

Teeth 250 2,400

Table 21.7.2_5

Mara Rosa Project

Wheel Tractor-Scraper and Tractor Wear Parts – Life Cycle and Costs

Part

Bulldozer Motor grader

Life Cycle Cost Life Cycle Cost

(h) (US$) (h) (US$)

Blade 350 4,800 350 1,900

Ripper 500 2,100 - -

The costs for wear parts for each piece of equipment were estimated in compliance with

manufacturer recommendations.



21.7.3 Equipment – Hourly Costs

Drilling Equipment

The productivity rate of the fragmentation drilling is based on an average penetration rate of

15m/h for the Rock Drill Rig (Ore) and 30 m/h for the Rock Drill Rig (Waste). Burden and

spacing grids of 3 m x 6 m and 3.5 m x 6.5 m were specified for the ore and rock waste,

respectively. The amount of sub-drilling for the 10 m benches is estimated at 0.7 m. During

the drilling operation, monitoring and analysis of the blasting results for each type of drilling

pattern that will be used is recommended in order to attain the best results.

The direct operating unit costs for fragmentation drilling by type of equipment are presented in

Tables 21.7.3_1 and 21.7.3_2 below (operator costs are not included). The hourly costs for

spare parts appear as a function of hourly equipment usage.



Table 21.7.3_1

Mara Rosa Project

Hourly Cost Details for Rock Drill Rig (Ore)

Rock Drill Rig

Ore Technical Information: Φ = 4”, 102 mm

ITEM Unit Cost Consumption Durability Performance Hourly Cost

Diesel 1.02 US$/L 60.00 L/h 61.41 US$/h

Lubricant 3.79 US$/L 0.3 L/h 1.14 US$/h

Drilling Parts

- Bit 1,686 US$ 1,000 m 15.0 m/h 25.29 US$/h

- Hammer 8,419 US$ 4,000 m 31.57 US$/h

- Crossover 860 US$ 8,000 m 1.61 US$/h

- Drill Tubes 1,743 US$ 8,000 m 3.27 US$/h

- Adapter 1,290 US$ 8,000 m 2.42 US$/h

- Rotation Unit 1,290 US$ 8,000 m 2.42 US$/h

Subtotal 129.12 US$/h

Other variables according to use US$/hr

Equipment Life Cycle (h) 0 – 3,400 3,400

– 6,800

6,800 –

10,200

10,200 –

13,600

13,600 –

17,000

17,000 –

20,400

20,400–

23,800

23,800 –

27,200 average

Replacement Parts 2.38 11.17 68.63 176.49 98.88 147.67 174.53 174.53 106.78

Consumables 129.12 129.12 129.12 129.12 129.12 129.12 129.12 129.12 129.12

Total Cost (US$/h) 131.50 140.29 197.75 305.61 228.00 276.79 303.65 303.65 235.91



Table 21.7.3_2

Mara Rosa Project

Hourly Cost Details for Rock Drill Rig (Waste)

Rock Drill Rig

Waste Technical Information: Φ = 5”, 127 mm

ITEM Unit Cost Consumption Durability Performance Hourly Cost

Diesel 1.02 US$/L 35.00 L/h 35.82 US$/h


Drilling Parts

- Bit 972 US$ 1,000 m 30.0 m/h 29.16 US$/h

- Collar 2,224 US$ 4,500 m 14.83 US$/h

- Drill Rod 1,467 US$ 1,500 m 29.33 US$/h




– 6,800

6,800 –

10,200

10,200 –

13,600

13,600 –

17,000

17,000 –

20,400

20,400–

23,800

23,800 –

27,200 average


Consumables 110.27 110.27 110.27 110.27 110.27 110.27 110.27 110.27 110.27

Total Cost (US$/h) 118.57 184.34 252.43 205.09 294.81 258.40 170.09 173.73 207.19



Loading

Hydraulic excavators will be used to perform the excavating and loading procedures. Coffey

Mining considered a 5.8 m³ hydraulic excavator to load the ore and a 10.5 m³ hydraulic

excavator to load the waste rock. Five cubic metre front loaders will be used as ancillary

loading equipment at the mine in case the hydraulic excavators are unavailable. They will also

be used for removing debris from operating areas and for recovering material from the ore

stockpile to maintain a constant material feed at the crushing circuit. The ore and waste

excavators have productivity rates of 1,246 t/h and 2,256 t/h, respectively. The direct unit

operational costs are provided in Tables 21.7.3_3, 21.7.3_4 and 21.7.3_5. They do not

include operator labour costs.



Table 21.7.3_3

Mara Rosa Project

Hourly Cost Details for the Hydraulic Excavator (Ore)

Hydraulic Excavator

Ore Technical Information (m3) 5.8

ITEM Unit Cost Consumption Durability Hourly Cost

Diesel 1.02 US$/L 62.00 L/h 63.46 US$/h


Excavation Parts

- Bucket 18,421 US$/ea 5,000 h/ea 3.68 US$/h

- Teeth “Dipper” 2,632 US$/ea 200 h/ea 13.16 US$/h

Sub total 82.19 US$/h

Hourly spares cost estimate as equipment ages


– 11,275

11,275 –

16,913

16,913 –

22,550

22,550 –

28,188

28,188 –

33,826

33,826 –

39,463

39,463 –

45,101 Average


Consumables 82.19 82.19 82.19 82.19 82.19 82.19 82.19 82.19 82.19 Total Cost (US$/h) 90.33 116.38 131.97 136.23 136.33 137.86 140.35 142.35 129.01



Table 21.7.3_4

Mara Rosa Project

Hourly Cost Details for the Hydraulic Excavator (Waste)

Hidraulic Excavator

Waste Technical Information (m3) 10.5


Diesel 1.02 US$/L 147.00 L/h 149.9 US$/h


Excavation Parts


- Frame 119,581 US$/ea 20,000 h/ea 6.0 US$/h

- Teeth “Dipper” 63,925 US$/ea 5,000 h/ea 12.8 US$/h




– 11,275

11,275 –

16,913

16,913 –

22,550

22,550 –

28,188

28,188 –

33,826

33,826 –

39,463

39,463 –

45,101 Average


Consumables 172.8 172.8 172.8 172.8 172.8 172.8 172.8 172.8 172.8

Total Cost (US$/h) 180.94 206.98 222.57 226.83 227.24 228.47 230.96 232.96 180.94



Table 21.7.3_5

Mara Rosa Project

Hourly Cost Details for the Wheel Loader

Wheel Loader

Technical Information (m3) 5


Diesel 1.02 US$/L 28.00 L/h 28.7 US$/h


Tires 73,684 US$/set 4,500 h 16.4 US$/h

Excavation Parts


- Teeth “Dipper” 2,368 US$/ea 20 h/ea 9.5 US$/h



Equipment Life Cycle (h) 0 - 5638 5,638

– 11,275

11,275 –

16,913

16,913–

22,550

22,550 –

28,188

28,188 –

33,826

33,826 –

39,463

39,463 –

45,101 average


Consumables 56.69 56.69 56.69 56.69 56.69 56.69 56.69 56.69 56.69

Total Cost (US$/h) 64.56 92.56 100.03 106.61 112.34 113.01 111.80 114.34 101.91



Transport

The transport and productivity costs vary year-to-year depending on changes that occur to the

pit geometry and, consequently, changes to the average transport distances. An average

hourly fuel consumption rate was estimated according to the characteristics and average

speed of the equipment. A more detailed year-to-year study of the pit transport profiles should

be conducted in order to arrive at a more precise estimate of the fuel consumption of the

trucks as they drive uphill under load, drive downhill empty and move on horizontal surfaces.

The direct operational costs with respect to trucks are provided in Table 21.7.3_6 and

21.7.3_7. They do not include operator labour costs.



Table 21.7.3_6

Mara Rosa Project

Hourly Cost Details for the Haul Truck (Ore)

Scania truck model G470 CB10x4 Technical

Information (t)

45


Diesel 1.02 US$/L 25.00 L/h 25.6 US$/h


Tires 12,632 US$/set 4,500 hr/set 2.8 US$/h




– 11,275

11,275 –

16,913

16,913 –

22,550

22,550 –

28,188

28,188 –

33,826

33,826 –

39,463

39,463 –

45,101

Average


Consumables 29.53 29.53 29.53 29.53 29.53 29.53 29.53 29.53 29.53

Total Cost (US$/h) 34.53 66.20 74.89 82.06 87.42 85.58 85.11 88.81 75.58



Table 21.7.3_7

Mara Rosa Project

Hourly Cost Details for the Haul Truck (Waste)

Haul Truck Technical

Information 95 t

Waste


Diesel 1.02 US$/L 87.00 L/h 89.0 US$/h


Tires 176,842 US$/set 5 000 hr/set 35.4 US$/h




– 11,275

11,275 –

16,913

16,913 –

22,550

22,550 –

28,188

28,188 –

33,826

33,826 –

39,463

39,463 –

45,101

Average


Consumables 127.8 127.8 127.8 127.8 127.8 127.8 127.8 127.8 127.8

Total Cost (US$/h) 138.19 139.40 175.05 181.54 183.03 187.15 186.95 188.85 172.52



Ancillary Equipment

The operational cost estimates and tabulations for the ancillary equipment are provided in

Fonseca and Horta (2011). Operator labour costs are not included.

21.7.4 Operating Cost by Activity

The cost for each activity was estimated by taking into account the hours that were expended

for each specific unitary operation and its respective hourly cost. The cost estimate for rock

blasting and fragmentation was reached in accordance with a proposal submitted by Britanite

IBQ, a contractor.

The yearly hourly costs were calculated according to how each piece of equipment was used.

Drilling

The drilling costs refer to the ROC D65 and ROC F9 diesel drill rigs, manufactured by Atlas

Copco. The operating costs, which are provided in Portuguese in the Mining Study Appendix,

do not include operator labour costs.

Blasting and Fragmentation

The blast supervisor of the blasting contractor will supervise the explosive charging, blasting

and fragmentation operations. Costs estimates are provided in a proposal that was submitted

by Britanite IBQ.

Tables 21.7.4_1, 21.7.4_2 and 21.7.4_3 present the blasting and fragmentation cost estimate

for ore and waste rock taking into account a mixture of 60% emulsion and 40% ANFO,

including the emulsion plant, explosives and labour costs for the contractor personnel.

Blasting will be performed on the 5m benches at the contact points between the ore and host

rock in order to provide a more selective extraction process. Plans stipulate that 30% of ore

blasting and fragmentation occur in 5m benches, while the remainder should occur in 10m

benches.

Table 21.7.4_1

Mara Rosa Project

Estimated Blasting and Fragmentation Unit Cost – 10m Bench

Unit Costs – 10m Bench – 4” – Ore

Explosives and Accessories Consumption Cost

Unit Cost (US$) (kg pcs)/m3 (US$/t)

Ibegel 4" x 24 $2.27 0.05 0.12

Anfomax $1.99 0.08 0.15

Brinel MF 12 m detonators $7.71 0.002 0.02

Total 0.29



Table 21.7.4_2

Mara Rosa Project

Estimated Blasting and Fragmentation Unit Cost – 5m Bench

Unit Costs – 5m Bench – 4” – Ore



Ibegel 4" x 24 $2.27 0.05 0.12

Anfomax $1.99 0.08 0.15


Total 0.30

Table 21.7.4_3

Mara Rosa Project

Estimated Blasting and Fragmentation Unit Cost – 10m Bench Unit Costs – 10m Bench – 5” – Waste Rock



Ibegel 4" x 24 $2.27 0.04 0.09

Anfomax $1.99 0.06 0.11


Total 0.21

Loading

The loading cost takes into account the total estimated number of hours and the hourly cost

for each piece of equipment that is performing the loading. The excavators that were selected

to remove ore were 5.8 m3 capacity, while those chosen to remove waste rock were 10.5 m3

capacity. Both models are of the backhoe variety. The type of excavator (backhoe or shovel)

should be chosen during the Feasibility Study. The front loader that was selected was of 5 m3

capacity.

Transport

The transport cost takes into account the total number of hours expended and the cost

estimate for each fleet of trucks. A 45 t capacity truck was selected for transporting ore. A 95 t

capacity truck will transport the waste rock.

21.7.5 Consumption of Diesel Fuel

The total fuel consumption per year is presented in Table 21.7.5_1. A total consumption of

38.34 x 106 litres of diesel fuel was estimated for the entire life of the project.



Table 21.7.5_1

Mara Rosa Project

Total Diesel Fuel Consumption per Year

Year 0 1 2 3 4 5 6 7

Diesel (l) 3,563,363 5,752,887 7,491,843 7,985,497 9,423,614 9,375,575 5,348,644 2,905,806

L/kt ore 17,788 2,503 3,116 3,258 3,862 3,724 2,189 1,231

21.7.6 Labour Cost and Requirement

Direct Labour at the Mine

The direct operating labour was estimated by taking into account the work schedule and the

amount of equipment that would be used during each year of mine operation. Four groups of

personnel have been considered in this study in which three groups work while one rests,

maintaining a rotation.

Drill rigs were considered as having one operator and an assistant per shift, while the other

equipment will be operated by one operator per shift.

Table 21.7.6_1 presents the number of mobile equipment necessary and the direct amount of

labour that is involved. Note that a maximum number of 144 staff members will be necessary.

Table 21.7.6_1

Mara Rosa Project

List of Operating Equipment and Labour

Mine Equipment Year

0 1 2 3 4 5 6 7

Main Equipment

Hydraulic Excavator (Ore) 1 2 2 2 2 2 2 2

Hydraulic Excavator (Waste) 1 2 2 2 2 2 2 2

Haul Truck (Ore) 1 2 2 3 3 3 3 3

Haul Truck (Waste) 4 6 9 10 13 13 6 2

Wheel Loader 2 2 2 2 2 2 2 2

Rock Drill Rig (Ore) 1 2 2 2 2 2 2 2

Rock Drill Rig (Waste) 2 3 3 3 3 3 2 2

Ancillary Equipment

Bulldozer 2 2 2 2 2 2 2 2

Motor Grader 1 1 1 1 1 1 1 1

Water Truck 1 1 1 1 1 1 1 1

Service Truck 2 2 2 2 2 2 2 2

Truck Loading Crane 1 1 1 1 1 1 1 1

Flat-bed Truck – 3 axles 1 1 1 1 1 1 1 1

Crane 1 1 1 1 1 1 1 1

Total 21 28 31 33 36 36 28 24

Labour 84 112 124 132 144 144 112 98



Mine and Support Staff

A maximum of 194 staff members in year 5 were estimated for managing and providing

support services at the mine. The reader should note that no personnel have been counted

for blasting operations on the assumption this activity will be contracted out. Labour costs for

blasting activities are captured in unit blasting costs. Table 21.7.6_2 presents the yearly mine

labour requirements.

Table 21.7.6_2

Mara Rosa Project

Mine Staff

Total Mine Staff Year

0 1 2 3 4 5 6 7

Mine Manager 1 1 1 1 1 1 1 1

Mine Planning Engineer 1 1 1 1 1 1 1 1

Mine Geologist 1 1 1 1 1 1 1 1

Administrative Assistant 1 1 1 1 1 1 1 1

Shift Supervisor 4 4 4 4 4 4 4 4

Drill rig Assistant 12 20 24 24 24 24 24 16

Mine Infrastructure Supervisor 1 1 1 1 1 1 1 1

Equipment Operators 84 112 124 132 144 144 112 98

Geology/Mining Technician 2 2 2 2 2 2 2 2

Geology Mine/Infrastructure Assistant 8 8 8 8 8 8 8 8

Topographer 1 1 1 1 1 1 1 1

Topography Assistants 2 2 2 2 2 2 2 2

Sampling Assistant 4 4 4 4 4 4 4 4

Total 122 158 174 182 194 194 162 140



Mine Maintenance and Warehouse Personnel

The direct labour requirement was estimated according to operation levels for each year. One

mechanic, one electrician and 2 assistants, per shift, are required for the main equipment.

Table 21.7.6_3 presents the maintenance and warehouse labour requirements by year.

Table 21.7.6_3

Mara Rosa Project

Mine Maintenance & Warehouse Personnel

Total Mine Maintenance & Warehouse Staff

Year

0 1 2 3 4 5 6 7

Maintenance

Supervisor (Maintenance Planning) 1 1 1 1 1 1 1 1

Supervisor (Electrical) 1 1 1 1 1 1 1 1

Supervisor (Mechanical) 1 1 1 1 1 1 1 1

Supervisor (Instrumentation and Automation) 1 1 1 1 1 1 1 1

Electrician 12 16 16 16 20 20 16 12

Mechanic 12 16 16 16 20 20 16 12

Pump operator 4 4 4 4 4 4 4 4

Assistants 20 28 32 32 36 36 28 24

Warehouse

Stockman 1 1 1 1 1 1 1 1

Fork-lift operator 1 1 1 1 1 1 1 1

Stockman Assistant 2 2 2 2 2 2 2 2

Total 56 72 76 76 88 88 72 60

Other costs, such as administrative, laboratory, first-aid station and outsourced service costs,

which have not been addressed in this section, were included in the process costs, in addition

to plant warehouse and maintenance costs.

Total Mine Labour Cost

An estimate for total year-to-year labour cost was prepared considering the total amount of

personnel per year and taking into account their salaries. The payroll taxes were also

included.

A provision for replacement workers to cover holidays and 10% absenteeism is included in

the personnel costs,

An annual uniform allowance of US$500 and a monthly transportation and food allowance of

R$180 was estimated for each employee.

Total payroll burdens represented 87% of the direct salaries.

Table 21.7.6_4 presents the base salaries and annual expenses for the labour requirements

of the mine.



Table 21.7.6_4

Mara Rosa Project

Total Mine Labour Cost

Sector Base Salary Labour Cost (US$/year) M.O Total

(US$) (US$/year) 0 1 2 3 4 5 6 7

Mine

Mine Manager 183,222 183,222 183,222 183,222 183,222 183,222 183,222 183,222 183,222 1,465,776

Mine Planning Engineer 139,050 139,050 139,050 139,050 139,050 139,050 139,050 139,050 139,050 1,112,400

Geologist 104,545 104,545 104,545 104,545 104,545 104,545 104,545 104,545 104,545 836,360

Administrative Assistant 25,618 25,618 25,618 25,618 25,618 25,618 25,618 25,618 25,618 204,944

Shift Supervisor 51,288 205,152 205,152 205,152 205,152 205,152 205,152 205,152 205,152 1,641,216

Drill Rig Assistant 12,926 155,112 258,520 310,224 310,224 310,224 310,224 310,224 206,816 2,171,568

Mine Infrastructure Supervisor 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 410,304

Equipment Operators 25,618 2,049,448 2,766,755 3,176,645 3,279,117 3,689,007 3,689,007 2,971,700 2,356,866 23,978,545

Geology/Mining Technician 51,288 102,576 102,576 102,576 102,576 102,576 102,576 102,576 102,576 820,608

Geology/Infrastructure Assistant 25,618 204,945 204,945 204,945 204,945 204,945 204,945 204,945 204,945 1,639,560

Topographer 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 410,304

Topography Assistant 25,618 51,236 51,236 51,236 51,236 51,236 51,236 51,236 51,236 409,888

Sampling Assistant 12,926 51,704 51,704 51,704 51,704 51,704 51,704 51,704 51,704 413,632

Subtotal 3,375,184 4,195,899 4,657,493 4,759,965 5,169,855 5,169,855 4,452,548 3,734,306 35,515,105

Maintenance

Supervisor (Maintenance Planning) 139,050 139,050 139,050 139,050 139,050 139,050 139,050 139,050 139,050 1,112,400

Supervisor (Electrical) 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 410,304

Supervisor (Mechanical) 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 410,304 Supervisor (Instrumentation and Automation)

51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 51,288 410,304

Electrician 25,618 307,417 409,890 409,890 409,890 512,362 512,362 409,890 307,417 3,279,118

Mechanic 25,618 307,417 409,890 409,890 409,890 512,362 512,362 409,890 307,417 3,279,118

Pump operator 16,545 66178 66178 66178 66178 66178 66178 66178 66178 529,424

Assistants 16,545 330,892 463,248 529,427 529,427 595,605 595,605 463,248 397,070 3,904,522

Subtotal 1,304,818 1,642,120 1,708,299 1,708,299 1,979,421 1,979,421 1,642,120 1,370,996 13,335,494

Warehouse

Stockman 34,210 34,210 34,210 34,210 34,210 34,210 34,210 34,210 34,210 273,680

Fork-lift operator 25,618 25,618 25,618 25,618 25,618 25,618 25,618 25,618 25,618 204,944

Stockman Assistant 12,925 25,851 25,851 25,851 25,851 25,851 25,851 25,851 25,851 206,808

Subtotal 85,679 85,679 85,679 85,679 85,679 85,679 85,679 85,679 685,432

Uniform, Meal and Transport for each person per year 462,840 601,160 665,000 675,640 750,120 750,120 633,080 516,040 5,054,000

Total 5,228,521 6,524,858 7,116,471 7,229,583 7,985,075 7,985,075 6,813,427 5,707,021 54,590,031



21.7.7 Operating Cost Summary

Table 21.7.7_1 summarizes each of the mine’s operating costs. The cost of the

Administration, Laboratory and First-Aid Station staff is included in the process plant cost,

which also includes the warehouse costs and the maintenance costs of the plant.



Table 21.7.7_1

Mara Rosa Project

Summary of the Total Operating Cost for the Mine

Year 0 1 2 3 4 5 6 7 TOTAL

Total Ore Mass (kton) 200 2,298 2,404 2,451 2,440 2,517 2,444 2,361 17,117

Total Rock Waste

Mass (kton) 11,003 18,139 24,324 24,579 23,660 22,961 9,790 2,733 137,188

Total Mass (kton) 11,203 20,438 26,729 27,030 26,100 25,478 12,233 5,094 154,305

Total Equipment

Total Cost (US$) 8,974,921 16,506,581 24,159,654 25,613,160 29,541,990 29,366,010 16,808,504 10,047,225 161,018,046

Unit Cost (US$/t Ore)

44.80 7.18 10.05 10.45 12.11 11.67 6.88 4.26 9.41

Unit Cost (US$/t Material)

0.80 0.81 0.90 0.95 1.13 1.15 1.37 1.97 1.04

Total Labour Force

Sub total (US$) 5,237,088 6,533,425 7,125,037 7,238,149 7,993,642 7,993,642 6,821,993 5,715,588 54,658,564


26.14 2.84 2.96 2.95 3.28 3.18 2.79 2.42 3.19


0.47 0.32 0.27 0.27 0.31 0.31 0.56 1.12 0.35

Total Equip. Mine + Labour Force

Total Cost (US$) 14,212,009 23,040,006 31,284,691 32,851,309 37,535,631 37,359,652 23,630,498 15,762,813 215,676,610


70.95 10.02 13.01 13.40 15.38 14.84 9.67 6.68 12.60


1.27 1.13 1.17 1.22 1.44 1.47 1.93 3.09 1.40



The average operating cost for mining, segmented by major cost areas, is presented in Table

21.7.7_2.

Table 21.7.7_2

Mara Rosa Project

Average Operating Cost

Item % US$/t material

Blasting and Fragmentation 16.32 0.230

Diesel fuel and Lubricants 28.26 0.400

Electricity 0.35 0.005

Replacement Parts 29.74 0.420

Labour 25.34 0.350

Total 100.00 1.405

An average mine operating cost of US$1.055/t of transported material was estimated for the

mine equipment. The average mine labour cost is US$0.35/t of transported material.

The average figure for the total mine operating cost is US$1.40/t of material, or US$12.59/t of

ore. This figure does not include contingencies and benefits. This unit cost is presented for

Years 0 to 7 inclusive. The reader should note that pre-stripping in Year 0 is capitalised in the

economic analysis; consequently a different operating cost is estimated and presented in

Section 22.

21.8 Plant Operating Costs

21.8.1 Basis

There will be applied a training scheme so that as much of the labour (with the exception of

management positions) to operate the mine and the plant will come from the surrounding

area. This approach has been successful in the past and this can be seen in the nearest mine

of Chapada (Mineração Maraca) owned by Yamana only 35 kilometres from the Posse mine.

The operating cost for contracting the entire plant personnel (including administration) has

been included for 2013. This means that during construction there will be a full complement of

Amarillo staff to help administer the project as well as allow time for training of local

candidates.

The salaries have been derived using as a base a recent industry study carried out by the

mining company Rio Novo whilst reagent consumption prices have all been based on at least

one firm supply offer. The consumption estimates used for calculating the quantities of

reagents used at the various stages have been taken from testwork or, in the case of mill ball

consumption, from experience (a total of 1,000 grams per tonne of ore has been used).

Electrical consumption has been calculated based on the electric motor power rating.

Table 21.8.1_1 shows the unit costs taken to calculate plant operating costs.



Table 21.8.1_1

Mara Rosa Project

Plant Unit Costs Item Unit Unit Cost $US

New Product

Hours

Variable Operating Costs

Feed Material

Electric Power kWh $0.037

Process water m3 $0.14

Process water- potable m3 $0.10

Sectional Mining

Ore and waste mining t

Sectional - Processing Plant

Crusher linings each $250,000.00

LPG L $0.79

Ball Mills Liners each $200,000.00

Mill Balls, primary mill kg $1.77

Mill Balls, secondary mill kg $2.08

Hydrochloric Acid, 33% commercial t $249.11

Hydrated Lime, includes transport t $194.74

Sodium cyanide, includes transport kg $3.50

Leach Aid kg $48.61

Sodium Metabisulphite kg $1.12

Copper Sulphate kg $5.50

Flocculant Kg $10.24

Caustic Soda kg $0.60

Laboratory reagents kg $10.00

Activated Carbon kg $2.59

Freight t $108.95

General Consumables $ $0.98

Regular Maintenance % Capital Unit costs have been derived using the following exchange rates as appropriate:

US$1 = AUD$1.02; R$1.90; CAD$1.03.

The required personnel to operate the plant are shown in Table 21.8.1_2 (the calculated cost

per hour is included in R$ and US$, using a factor of 1.8776 to account for holidays/13th

salary and payroll tax).



Table 21.8.1_2

Mara Rosa Project

Plant Personnel Requirements and Cost Estimate

Area Function Day Shift 1 Shift 2 Shift 3 Annual Salary Monthly Annual Cost * R$/hr US$/person US$/hr

Process Plant Manager 1 185,409 15,451 348,124 167 183,223 88

Senior Metallurgist 1 105,792 8,816 198,635 95 104,545 50

Junior Metallurgist 1 92,111 7,676 172,948 83 91,025 44

Metallurgical Clerk 1 34,618 2,885 64,999 31 34,210 16

Shift Supervisors 2 1 1 1 51,900 4,325 97,447 47 51,288 25

Crusher Operator 1 1 1 1 25,924 2,160 48,675 23 25,618 12

Control Room Operator 1 1 1 1 34,618 2,885 64,999 31 34,210 16

Mill Operator 1 1 1 1 25,924 2,160 48,675 23 25,618 12

Leach Operator 1 1 1 1 25,924 2,160 48,675 23 25,618 12

Elution and Gravity Sep. Operator 1 1 1 1 25,924 2,160 48,675 23 25,618 12

Gold Room Operator 1 34,618 2,885 64,999 31 34,210 16

Assistent Crusher Operator 1 1 1 1 16,742 1,395 31,435 15 16,545 8

Reagents and Infrastructure 6 13,080 1,090 24,559 12 12,926 6

Sub-total 19 7 7 7

Maintenance Maintenence Manager 1 185,409 15,451 348,124 167 183,223 88

Maintenence Planner 1 140,709 11,726 264,195 127 139,050 67

Electrical Supervisor 1 51,900 4,325 97,447 47 51,288 25

Mechanical Supervisor 1 51,900 4,325 97,447 47 51,288 25

Instrumentation Supervisor 1 51,900 4,325 97,447 47 51,288 25

Mechanics 4 2 25,924 2,160 48,675 23 25,618 12

Electricians 4 2 25,924 2,160 48,675 23 25,618 12

Munck Operator 1 25,924 2,160 48,675 23 25,618 12

Greaser 1 25,924 2,160 48,675 23 25,618 12

Sub-total 15 4 0 0

Stores Storeman 1 34,618 2,885 64,999 31 34,210 16

Fork-lift Operator 1 25,924 2,160 48,675 23 25,618 12

Storeroom Assistents 2 13,080 1,090 24,559 12 12,926 6

Sub-total 4 0 0 0

Laboratório Chemist 1 105,792 8,816 198,635 95 104,545 50



Area Function Day Shift 1 Shift 2 Shift 3 Annual Salary Monthly Annual Cost * R$/hr US$/person US$/hr

Technicians (chemical) 2 51,900 4,325 97,447 47 51,288 25

Sample Preparation 2 2 2 2 13,080 1,090 24,559 12 12,926 6

Sub-total 5 2 2 2

Ambulance Nurse 1 1 1 1 51,900 4,325 97,447 47 51,288 25

Administration General Manager 1 288,161 24,013 54,1051 260 284,764 137

Administrative Manager 1 140,749 1,1729 26,4270 127 139,090 67

Financial Controller 1 140,749 1,1729 26,4270 127 139,090 67

Secretary 1 51,900 4,325 97,447 47 51,288 25

Accounting Clerks 2 34,618 2,885 64,999 31 34,210 16

Telephonist 1 13,080 1,090 24,559 12 12,926 6

Chief buyer 1 120,968 10,081 22,7130 109 11,9542 57

Assistant Buyers 2 34,618 2,885 64,999 31 34,210 16

Safety Officer 1 120,968 10,081 22,7130 109 119,542 57

Technician (safety) 1 51,900 4,325 97,447 47 51,288 25

Environmental Officer 1 120,968 10,081 22,7130 109 119,542 57

Technician, (environment) 1 51,900 4,325 97,447 47 51,288 25

Technician (Information technology) 1 92,100 7,675 172,927 83 91,014 44

Manager HR 1 185,409 15,451 348,124 167 183,223 88

Assistant (Human Resources) 1 51,900 4,325 97,447 47 51,288 25

Drivers 3 25,924 2,160 48,675 23 25,618 12

Sub-total 20 0 0 0 0 0 0 0 0

Grand Totals 64 14 10 10 0 0 0 0 0

Total Final 98 0 0 0 0 0

Services Contracted 0 0 0 0 0

(Number of people estimated) 0 0 0 0 0

Cleaning, bus drivers 8 13,080 1,090 24,559 12 12,9267 6

Security, Kitchens, 32 13,080 1,090 24,559 12 12,926 6

The above numbers do not include US$200/month, estimated cost of meals and daily travel .

Hours per annum = 2080

*Cost factor, (payroll tax) 1.8776.

Using US$1 = R$1.9.



Table 21.8.1_3 presents the salaries of the various employee categories expressed in US$/h.

Table 21.8.1_3

Mara Rosa Project

Staff Salaries (Admin and Plant)

Direct Labour per hour including taxes US$/h

ADMIN

General Managers 136.84

Admin. Manager 66.84

Financial Manager 66.84

Executive Secretary 24.74

Accounting Clerks 16.32

Telephone Operator 6.32

Chief buyer 57.37

Assistant Buyers 16.32

Safety Engineer 57.37

Safety Technician 24.74

Environmental Engineer 57.37

Environmental Technician 24.74

Information Technician 43.68

Manager Human Resources 87.89

Human Resources Assistant 24.74

Drivers 12.11

Storeman 16.32

Fork-lift Operator 12.11

Assistant Storeman 6.32

Nurses 24.74

Plant

Plant Manager 87.89

Senior Met. 50.00

Junior Met 43.68

Metallurgical Clerk 16.32

Shift Supervisors 24.74

Crusher Operator 12.11

Assistant Crusher Operator 7.89

Control Room Operator 16.32

Mill Operator 12.11

Leach Operator 12.11

Elution and Grav. Sep. Operator 12.11

Gold Room Operator 16.32

Infrastructure and Reagents 6.32

Laboratory

Chief Chemist 50.00

Chemical Technicians 24.74

Samplers 6.32

Plant Maintenance

Maintenance Manager 87.89

Maintenance Planner 66.84

Electrical Supervisor 24.74

Mechanical Supervisor 24.74

Instrument Supervisor 24.74

Mechanics/Welders, etc 12.11

Electricians 12.11

Munck Operator 12.11

Lubricator 12.11

Contract Maintenance 23.68

Reg Contract - Cleaning/Security 6.32

Meals and local transport/person/month 178.95



21.8.2 Overall Plant Operating Costs

Plant Operating Costs

Table 21.8.2_1

Mara Rosa Project

Plant Operating Cost Estimate

Units Usage/t Hourly $/t %

Electric Power kWh 35.8 11336.3 1.33 11.52%

Process water m3 1.0 317.1 0.14 1.21%

Process water- potable m3 0.3 86.6 0.03 0.27%


Crusher linings each 0.30 2.59%

LPG L 95.1 0.24 2.05%

Ball Mills Liners each 0.32 2.77%

Total Reagents 4.46 38.54%

Freight t 0.6 0.22 1.88%

General Consumables $ 0.1 31.7 0.10 0.85%

Regular Maintenance % Capital 1.35 11.66%

Total Personel 1.87 16.17%

Head Office Costs 0.20 1.70%

Contract Maint 1.0 0.07 0.65%

Reg Contract - Cleaning/Security 40.0 0.21 1.82%

Meals and local transport/person/month 150.0 0.13 1.11%

Office Expenses 0.03 0.23%

Insurances 0.03 0.22%

Contingency (5%) 0.55 4.76%

11.56 100.00% The unit operating cost of US$11.56 in this table includes G&A

Table 21.8.2_1 is based on a throughput of 2.5 Mtpa. It includes a contingency of 5%. The

variable part of these costs can be considered as US$9.11 /t of ore with a value of US$2.45/t

as a total for fixed costs (which include all G & A costs).

G & A Costs

An annual value of US$480,000 was estimated as head office costs. The balance of

estimated G&A costs are as local costs and include an additional 5% contingency. The

estimation of G & A costs is presented in Table 21.8.2_2.



Table 21.8.2_2

Mara Rosa Project

G & A Cost Estimate

Units Usage/t Hourly $/t %

Electric Power kWh 0.5 158.5 0.02 1.02%

Process water m3

Process water- potable m3 0.1 15.9 0.01 0.31%


Crusher linings each

LPG l

Ball Mills Liners each

Total Reagents

Freight t 0.0 0.6 0.22 11.88%

General Consumables $ 0.1 31.7 0.10 5.34%

Regular Maintenance % Capital 0.03 1.64%

Total Personnel 0.79 42.90%

Head Office Costs 0.20 10.68%

Contract Maintenance 1.0 1.0 0.07 4.07%

Reg Contract - Cleaning/Security 40.0 0.21 11.46%

Meals and local transport/person/month 70.0 0.06 3.28%

Office Expenses 0.03 1.38%

Insurances 0.02 1.29%

Contingency 0.09 4.76%

1.83 100.00%

21.9 Summary of Operating Costs

Peak labour requirements are about 380 staff in Year 5, with 320 operations staff and an

estimated 40 contract staff. Coffey Mining and ONIX estimate that roughly two-thirds or 250

staff and contractors will be on site at any time.

Summary of Mine Operating Costs

Total mine operating cost is estimated as US$1.40/t of material. With a waste strip ratio of

8:1, this implies a mine operating cost of US$12.59/t of ore, including pre-stripping costs.

In the Economic Analysis (Section 22) pre-stripping costs have been capitalised, thus

reducing the mine operating unit costs during the production Years 1 to 7 to US$11.85/t of ore

(US$8.93/t of ore - variable costs and US$2.92/t of ore - fixed costs)

Summary of Plant Operating Costs

G&A costs are estimated to be US$1.83/t of ore treated whilst plant costs are calculated to be

US$9.73/t of ore processed.

These costs were estimated on the basis of 2.5 Mtpa plant through-put; the mining schedule

shows small deviations from this production rate that impact the plant operating and G & A

costs each year.



The reader should note that in the Economic Analysis (Section 22) the plant operating costs

have been re-calculated to adjust for the impact of fluctuations in annual production rates from

the mine pit. These costs are US$9.78 /t G&A costs and US$1.90 /t of ore processed

respectively



22 ECONOMIC ANALYSIS

The overall economics of the Mara Rosa Project have been evaluated using conventional

discounted cash flow techniques based on the production schedules, capital expenditures and

operating costs discussed in this report. This text should be read in conjunction with the Life

of Mine cash flow analysis in Section 23.2 below. In addition, the following key parameters

were integral to the cash flow model and the economic results:

The base case metal price was $1,200/oz gold;

The analysis is based on 100% equity financing with no debt component;

All costs and revenues are reported in “real” or constant US dollars without escalation;

An income tax rate of 25% was applied, based on the general understanding of Brazilian

income tax laws;

Provision was also made for the Brazilian Social Contribution Tax of 9%.

22.1 Cash Flow Assumptions

22.1.1 Mine Production Sequence

The mine production is based on the mine plan and sequencing described earlier in this

report. Approximately 17.1 Mt of ore will be mined over a period of seven years following a

one year construction and pre-production period. Based on a mine dilution of 3% and a mine

recovery of 97%, the average ROM gold grade is expected to be 1.72 g/t for an estimated

total of 945,208 oz of contained ROM gold.

22.1.2 Metallurgical Recovery

A metallurgical recovery of 92% has been assumed. The design capacity of the mill is

modeled at 2.5 Mtpa. An average of 123,374 oz of gold are expected to be produced on an

annual basis. The total gold production through the life-of-mine is estimated at 869,600 oz.

22.1.3 Metal Prices and Net Revenues

The long-term gold price incorporated into the cash flow analysis is US$1,200/oz. This price is

based on the 12-quarter moving average historical gold as shown in the following table. The

use of historical metal prices is a generally accepted methodology for modeling long-term

commodity prices.



Table 22.1.3_1

Mara Rosa Project

Average Historical Gold Prices

Average Gold US$/oz Au

Average Maximum Minimum

36-month 1,205 1,772 761

48-month 1,122 1,772 761

60-month 1,031 1,772 628

Basis: London Fix, October to October Maximum and minimum are based on monthly average values, not daily values.

Gross revenues are determined as the product of the recovered ounces of gold and the long-

term metal price. The gross revenues are adjusted to account for royalty charges and refining

costs to arrive at a net revenue value. These and other parameters shown in Table 22.1.3_2

were applied in determining the annual net revenue values.

Table 22.1.3_2

Mara Rosa Project

Adjustments to Gross Revenues

Parameter Description Units Factor

Refining, transportation, insurance and sales

Assumed value % of gross revenues 1.5%

COFINS/PIS Federal tax similar to VAT – exports are exempt

% of gross revenues 0%

CFEM Federal compensation for exploitation of mineral resources. Value is dependent on mineral type.

% of gross revenues 1.0%

Land owners royalty Charged by federal government. Equivalent to 50% of CEFEM.

% of gross revenues 0.5%

Royalty Franco-Nevada Corporation % of gross revenues 1.0%

Royalty Royal Gold, Inc. % of gross revenues 1.0%



22.1.4 Operating Costs

The operating costs are discussed in detail in Section 21 of this report. The average

operating costs incorporated into the cash flow analysis are shown in Table 22.1.4_1.

Table 22.1.4_1

Mara Rosa Project

Average Unit Production Costs Full Production Years

Cost Item Description Units Average1

Mining Operating Costs

Variable Costs Includes drilling, stripping, excavation, loading, transportation, dewatering

US$/t 8.93

Fixed Costs Includes labour US$/t 2.92

Total US$/t 11.85

Processing Operating Costs2

Variable Costs Includes power, reagents and consumables, regular maintenance

US$/t 8.11

Fixed Costs Includes labour, general and administrative, contingency

US$/t 3.57

Total US$/t 11.68

Summary

Mining and Processing US$/t 23.53

Cash Costs (exclusive of royalties and refining charges) US$/oz 464

Cash Costs (inclusive of royalties and refining charges) US$/oz 524 1 Average of costs in full production years. 2 The average processing costs are determined on the basis of the design unit cost of US$11.56/t at the design annual feed rate of 2.5 Mtpa.

22.1.5 Capital Expenditures, Depreciation and Amortization

The capital costs included in the cash flow analysis are summarized in the Table 22.1.5_1.

Table 22.1.5_1

Mara Rosa Project

Capital Cost Summary

Capital Cost Item Initial US$ M Sustaining US$ M Total US$ M

Mine Development (pre-strip) 14.2 14.2

Mining Equipment 24.6 9.7 34.4

Processing (inc. TSF) 91.4 9.6 100.9

Infrastructure 13.9 13.9

Studies and Management 14.0 14.0

Miscellaneous 3.4 3.4

Rebate from CLEG (7.9) (7.9)

Insurance 0.4 0.4

Process Contingency 16.2 16.2

Initial Working Capital 5.5 5.5

Total Capital 183.6 11.4 195.0 Rounding has been applied



Both the capital expenditures and the development costs have been fully depreciation or

amortized on a unit production basis over the life of the mine. Based on Brazilian rules, if the

useful life of an asset is longer than the concession period of the mine, a unit production

depreciation method can be applied.

Development costs include pre-stripping, excavation and civil works, tailings dam

construction, studies and management, insurance and other owner costs. These

development costs total US$60 M and are included in the various cost categories in the

capital cost summary shown in Table 22.1.5_1.

22.1.6 Salvage Value

A salvage value of US$23 million has been included in the cash flow for year 8. This value is

based on general assumptions regarding the recoverable salvage value of the mining and

processing equipment, spares and structural and systems equipment. Factors ranging

between 10% and 40% were applied to the original capital cost in determining the salvage

value. These factors were determined on the basis of general experience with other projects.

22.1.7 Taxes

The standard Brazilian corporate income tax rate is 15% which is increased by a surtax of

10% on taxable profits that exceed R$240,000 annually, resulting in a total income tax rate of

25%. In addition, a Social Contribution Tax of 9% is levied on the income before taxes. In

essence, the overall tax rate is 34%.

Brazilian tax laws allow for tax losses to be carried forward indefinitely. Tax losses are

allowed to offset up to 30% of the taxable income in any year.

22.1.8 Working Capital

An initial working capital allowance of US$5.5 M has been included in year 0. This amount

represents approximately 2 months of operating cost requirements.

The annual calculated cash flows were reconciled to account for annual changes in accounts

receivable and accounts payable. The basis for the reconciliation is a one month balance for

the accounts receivable (based on revenues), a two month balance for accounts payable as

based on reagents and consumables and a one month balance for accounts payable based

on labour costs. The initial working capital and outstanding accounts receivable and accounts

payable balances are recovered in year 8.

22.1.9 Closure Costs

Closure costs of US$8.5M were included in year 8. The most significant closure cost is

expected to be associated with the TSF. Section 16.2.5 provided an outline of the tailings

rehabilitation work that will be required. Material quantities have also been estimated in

Saunders (2011) and these form the basis for a preliminary cost estimate that will be

evaluated in more detail as part of a feasibility study.



22.2 Financial Performance

Under the estimates and assumptions used for the base cash flow analysis, the Mara Rosa

Project would be expected to generate an undiscounted, life-of-mine after tax cash flow of

US$273.7 M. At a discount rate of 5% the after tax NPV is US$178.5 M. The life-of-mine

economics are presented in Table 22.2_1.

Table 22.2_1

Mara Rosa Project

Life-of-mine Economics (US$)

Tonnes of Ore Processed (000s) 17,117

Average ROM Grade, g/t Au 1.72

Gold Ounces Sold (000s) 869,592

Total Revenues (M) 1,044

Revenue per tonne 61.29

Mining Cost per tonne (Year 1 to 7) * 11.85

Processing Cost per tonne (at design 2.5 Mtpa) 9.73

G&A Cost per tonne (at design 2.5 Mtpa) 1.83

Processing Cost per tonne (at scheduled plant throughput) 9.78

G&A Cost per tonne (at scheduled plant throughput) 1.90

Operating Cost per ounce 464

Operating Cost per ounce (including refining and Royalties) 524

Capital Costs (millions) 189.5

Initial Working Capital (millions) 5.5





Internal Rate of Return (after tax) 26.6%

Payback Period (after tax, production years) 3.0 * Note that mining cost, including pre-stripping in Year 0, is US$12.59 per tonne

The life of mine cashflow is presented in Table 22.2_2.



Table 22.2_2 Mara Rosa Project

Life-of-mine CashFlow

Am a rillo Gold Corpora tionP os s e De pos it, M a ra Ros a , Goiá s S ta te , Bra zilCa s h Flow M ode l - S a le sDRAFT - W ork in P rogre s s

U S $ (0 0 0 s )

P e r iod End Ye a r 0 Ye a r 1 Ye a r 2 Ye a r 3 Ye a r 4 Ye a r 5 Ye a r 6 Ye a r 7 Ye a r 8Ye a r Tota l/Avg 2 0 1 3 2 0 1 4 2 0 1 5 2 0 1 6 2 0 1 7 2 0 1 8 2 0 1 9 2 0 2 0 2 0 2 1

S ce n a rio C a p e x - O th e r 0 .0 0 D e p re c ia tio n /Am o riza tio n C a lcu la tio n Me th o d

Ke y P roduc tion P a ra m e te rsO re m in e d a n d m ille d k to n n e s 1 7 ,1 1 7 2 0 0 2 ,2 9 8 2 ,4 0 4 2 ,4 5 1 2 ,4 4 0 2 ,5 1 7 2 ,4 4 4 2 ,3 6 1W a s te d m in e d kto n n e s 1 3 7 ,1 8 8 1 1 ,0 0 3 1 8 ,1 3 9 2 4 ,3 2 4 2 4 ,5 7 9 2 3 ,6 6 0 2 2 ,9 6 1 9 ,7 9 0 2 ,7 3 3

To ta l k to n n e s 1 5 4 ,3 0 5 1 1 ,2 0 3 2 0 ,4 3 8 2 6 ,7 2 9 2 7 ,0 3 0 2 6 ,1 0 0 2 5 ,4 7 8 1 2 ,2 3 3 5 ,0 9 4

G o ldR O M o re g ra d e g /t 1 .7 2 1 .0 1 1 .6 2 1 .8 8 1 .8 5 1 .6 9 1 .5 4 1 .6 4 1 .8 7C o n ta in e d R O M g o ld ko z 9 4 5 6 1 1 9 1 4 5 1 4 6 1 3 3 1 2 4 1 2 9 1 4 2Mill re co ve ry % 9 2 .0 % 9 2 .0 % 9 2 .0 % 9 2 .0 % 9 2 .0 % 9 2 .0 % 9 2 .0 % 9 2 .0 % 9 2 .0 %R e co ve re d G o ld ko z 8 7 0 6 1 1 0 1 3 4 1 3 4 1 2 2 1 1 5 1 1 9 1 3 0

Re ve nue R e co ve re d g o ld ko z 8 7 0 6 1 1 0 1 3 4 1 3 4 1 2 2 1 1 5 1 1 9 1 3 0Me ta l p rice U S $ /o z 1 ,2 0 0 1 ,2 0 0 1 ,2 0 0 1 ,2 0 0 1 ,2 0 0 1 ,2 0 0 1 ,2 0 0 1 ,2 0 0 1 ,2 0 0G ro s s re ve n u e s U S $ (0 0 0 s ) 1 ,0 4 3 ,5 1 1 7 ,1 6 7 1 3 1 ,8 4 6 1 6 0 ,6 0 1 1 6 0 ,8 3 8 1 4 6 ,7 8 3 1 3 7 ,4 2 0 1 4 2 ,4 8 8 1 5 6 ,3 6 7

R e fin in g , tra n s p o rta tio n , in s u ra n ce a n d s a le s % o f re ve n u e s 1 .5 0 % 1 .5 0 % 1 .5 0 % 1 .5 0 % 1 .5 0 % 1 .5 0 % 1 .5 0 % 1 .5 0 % 1 .5 0 %P IS % o f re ve n u e s 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 %C O FIN S % o f re ve n u e s 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 % 0 .0 0 %C E FE M % o f re ve n u e s 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 %L a n d o w n e rs ro ya lty % o f re ve n u e s 0 .5 0 % 0 .5 0 % 0 .5 0 % 0 .5 0 % 0 .5 0 % 0 .5 0 % 0 .5 0 % 0 .5 0 % 0 .5 0 %R o ya lty - Fra n co N e va d a % o f re ve n u e s 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 %R o ya lty - R o ya l G o ld % o f re ve n u e s 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 % 1 .0 0 %

Unit Cos ts - P roduc tion Ye a rsMin e u n it co s ts

Min e va ria b le co s ts - e q u ip m e n t co s ts U S $ /to n n e o re 8 .9 3 7 .1 7 1 0 .0 3 1 0 .4 3 1 2 .0 9 1 1 .6 5 6 .8 7 4 .2 6Min e fixe d co s ts - la b o u r U S $ /to n n e o re 2 .9 2 2 .8 4 2 .9 6 2 .9 5 3 .2 8 3 .1 8 2 .7 9 2 .4 2

To ta l m in e u n it co s ts U S $ /to n n e o re 1 1 .8 5 1 0 .0 1 1 2 .9 9 1 3 .3 8 1 5 .3 7 1 4 .8 3 9 .6 7 6 .6 8

P ro ce s s in g u n it co s tsP ro ce s s in g va ria b le co s ts U S $ /to n n e o re 8 .1 1 8 .1 1 8 .1 1 8 .1 1 8 .1 1 8 .1 1 8 .1 1 8 .1 1P ro ce s s in g fixe d co s ts U S $ /to n n e o re 1 .6 7 1 .7 6 1 .6 8 1 .6 5 1 .6 5 1 .6 0 1 .6 5 1 .7 1

To ta l p ro ce s s in g u n it co s ts U S $ /to n n e o re 9 .7 8 9 .8 7 9 .7 9 9 .7 6 9 .7 6 9 .7 1 9 .7 6 9 .8 2

G & A u n it co s tsG & A fixe d co s ts U S $ /to n n e o re 1 .9 0 2 .0 0 1 .9 1 1 .8 7 1 .8 8 1 .8 2 1 .8 8 1 .9 4

To ta l u n it o p e ra tin g co s ts U S $ /to n n e o re 2 1 .6 3 2 1 .8 7 2 4 .6 9 2 5 .0 1 2 7 .0 1 2 6 .3 6 2 1 .3 0 1 8 .4 4U S $ /o z g o ld 4 6 4 .1 8 4 5 7 .5 0 4 4 3 .5 3 4 5 7 .4 0 5 3 8 .8 2 5 7 9 .4 7 4 3 8 .4 5 3 3 4 .1 2

R e fin in g , tra n s p o rta tio n , in s u ra n ce a n d s a le s U S $ /o z g o ld 1 8 .0 0 1 8 .0 0 1 8 .0 0 1 8 .0 0 1 8 .0 0 1 8 .0 0 1 8 .0 0 1 8 .0 0R o ya ltie s U S $ /o z g o ld 4 2 .0 0 4 2 .0 0 4 2 .0 0 4 2 .0 0 4 2 .0 0 4 2 .0 0 4 2 .0 0 4 2 .0 0C a s h co s ts p e r o u n ce o f g o ld U S $ /o z g o ld 5 2 4 .1 8 5 1 7 .5 0 5 0 3 .5 3 5 1 7 .4 0 5 9 8 .8 2 6 3 9 .4 7 4 9 8 .4 5 3 9 4 .1 2

U n it P ro d u ctio n B a s is

R e tu rn to C o n te n ts



Net IncomeGross revenues US$(000s) 1,043,511 7,167 131,846 160,601 160,838 146,783 137,420 142,488 156,367 -Refining, transportation, insurance and sales US$(000s) (15,653) (108) (1,978) (2,409) (2,413) (2,202) (2,061) (2,137) (2,346) -Royalties US$(000s) (36,523) (251) (4,615) (5,621) (5,629) (5,137) (4,810) (4,987) (5,473) -

Net revenues US$(000s) 991,335 6,809 125,254 152,571 152,796 139,444 130,549 135,364 148,548 -

Production Costs

Mining costs US$(000s) (201,258) - (23,006) (31,239) (32,805) (37,496) (37,322) (23,622) (15,768) -

Processing costs US$(000s) (171,098) (5,661) (22,675) (23,535) (23,914) (23,826) (24,451) (23,853) (23,184) -

G&A costs (36,689) (4,586) (4,586) (4,586) (4,586) (4,586) (4,586) (4,586) (4,586) -

EBITDA US$(000s) 582,290 (3,438) 74,987 93,211 91,491 73,536 64,190 83,302 105,011 -

Depreciation US$(000s) (129,579) (823) (15,575) (19,213) (20,317) (18,541) (17,359) (17,999) (19,752) -Amortization US$(000s) (59,946) (400) (7,032) (8,663) (8,106) (8,198) (8,677) (8,997) (9,873) -

EBIT US$(000s) 392,765 (4,661) 52,380 65,335 63,067 46,796 38,155 56,307 75,386 -

Income taxes US$(000s) (98,191) - (11,930) (16,334) (15,767) (11,699) (9,539) (14,077) (18,847) -Social contribution taxes US$(000s) (35,349) - (4,295) (5,880) (5,676) (4,212) (3,434) (5,068) (6,785) -

Net Income from Operations US$(000s) 259,225 (4,661) 36,155 43,121 41,625 30,886 25,182 37,162 49,755 -

Pre Tax Cash FlowEBITDA US$(000s) 582,290 (3,438) 74,987 93,211 91,491 73,536 64,190 83,302 105,011 -Less: Capital expenditures US$(000s) (129,579) (119,834) (3,416) (1,358) (4,971) - - - - -Less: Development costs US$(000s) (59,946) (58,305) 2,632 (547) 2,632 (3,179) (3,179) - - -Less: Initial working capital US$(000s) - (5,475) - - - - - - - 5,475Less: Closure costs US$(000s) (8,500) - - - - - - - - (8,500)Less: Increases in accounts receivable US$(000s) - - (10,438) (2,276) (19) 1,113 741 (401) (1,099) 12,379Plus: Increases in accounts payable US$(000s) - - 3,736 169 62 51 87 (181) (186) (3,739)Plus: Proceeds from equipment sales US$(000s) 23,000 - - - - - - - - 23,000Project cash flow US$(000s) 407,265 (187,052) 67,501 89,199 89,194 71,520 61,840 82,720 103,727 28,615Cumulative project cash flow (187,052) (119,550) (30,351) 58,843 130,363 192,203 274,923 378,650 407,265

Internal Rate of Return % 37.9%NPV 0.0% Discount Rate US$(000s) 407,265

5.0% Discount Rate US$(000s) 283,1387.0% Discount Rate US$(000s) 244,650

10.0% Discount Rate US$(000s) 195,966

Payback period production years 2.34 N/A N/A N/A 2.34 N/A N/A N/A N/A N/A

After Tax Cash FlowNet Income from Operations US$(000s) 259,225 (4,661) 36,155 43,121 41,625 30,886 25,182 37,162 49,755 -Plus: Depreciation US$(000s) 129,579 823 15,575 19,213 20,317 18,541 17,359 17,999 19,752 -Plus: Amortization US$(000s) 59,946 400 7,032 8,663 8,106 8,198 8,677 8,997 9,873 -Less: Capital expenditures US$(000s) (129,579) (119,834) (3,416) (1,358) (4,971) - - - - -Less: Development costs US$(000s) (59,946) (58,305) 2,632 (547) 2,632 (3,179) (3,179) - - -Less: Initial working capital US$(000s) - (5,475) - - - - - - - 5,475Less: Closure costs US$(000s) (8,500) - - - - - - - - (8,500)Less: Increases in accounts receivable US$(000s) - - (10,438) (2,276) (19) 1,113 741 (401) (1,099) 12,379Plus: Increases in accounts payable US$(000s) - - 3,736 169 62 51 87 (181) (186) (3,739)Plus: Proceeds from equipment sales US$(000s) 23,000 - - - - - - - - 23,000Project cash flow US$(000s) 273,725 (187,052) 51,277 66,985 67,751 55,610 48,867 63,576 78,095 28,615Cumulative project cash flow (187,052) (135,775) (68,789) (1,038) 54,572 103,439 167,014 245,110 273,725

Internal Rate of Return % 26.6%NPV 0.0% Discount Rate US$(000s) 273,725

5.0% Discount Rate US$(000s) 178,4927.0% Discount Rate US$(000s) 149,160

10.0% Discount Rate US$(000s) 112,238

Payback period production years 3.02 N/A N/A N/A N/A 3.02 N/A N/A N/A N/A



22.2.1 Sensitivity Analysis

As shown in Table 22.2_1 and Figure 22.2_1, the economics of the Mara Rosa Project are

most sensitive to variation in gold price and less sensitive to variations in capital expenditures

and operating costs. Variations in feed grades would be expected to yield similar results to

variations in the gold price.

Table 22.2_1

Mara Rosa Project

Cash Flow Sensitivity After Tax NPV at 5% Discount Rate, US$M

30%

Decrease 20%

Decrease 10%

Decrease Base Case

10% Increase

20% Increase

30% Increase

Metal Price 24.7 76.0 127.2 178.5 229.7 281.0 332.2

Operating Costs 242.5 221.2 199.8 178.5 157.1 135.8 114.5

Capital Costs 217.5 204.5 191.5 178.5 165.5 152.5 139.5

Figure 22.2_1

Sensitivity Spider Diagram

0.0

50.0

100.0

150.0

200.0

250.0

300.0

350.0

‐30% ‐20% ‐10% 0% 10% 20% 30%

NPV, U

S$ m

illions

% Change in Input

After Tax NPV (5%) Sensitivity Analysis

Gold Price Operating Cost Capital Expenditures



23 ADJACENT PROPERTIES

Several significant mineral deposits occur in the Mara Rosa region including the Posse gold

deposit, the Zacarias gold-silver-barite deposit and the Chapada copper-gold deposit, in

addition to numerous historic prospects and garimpos.

A detailed discussion of these various properties was provided in Hoogvliet Contract Services

& Australian Exploration Field Services Pty Ltd (2011) that is filed at Sedar.com and to which

the reader is referred.



24 OTHER RELEVANT DATA AND INFORMATION

Nothing to report in this section.



25 INTERPRETATION AND CONCLUSIONS

25.1 Geology

The Posse Deposit is hosted by a mylonitic shear hosted zone in a high greenschist to low

amphibolite metamorphic terrain. The ore body strikes NE-SW and dips about 45° to the NW.

On average, the ore body is about 30m wide. Alteration is dominated by silicification,

sericitization, K-feldspar flooding and pyritization. Gold is positively correlated with the

intensity of silicification and total sulphide content, and occurs as 10 to 100 µm sized particles

along the margins of silicates and in association with pyrite and frohbergite.

The updated resource reported in July 2011 includes Inferred, Indicated and, for the first time,

Measured Mineral Resources. The 2011 drilling has significantly increased the confidence in

the resource estimate with the Inferred category being reduced from 42% of total resources in

the 2010 estimate to 12% in the 2011 estimate.

A total Measured and Indicated Mineral resource of 20.86 Mt at a grade of 1.75 g/t gold is

reported, together with an Inferred Mineral Resource of 3.63 Mt at a grade of 1.34 g/t gold.

The opinion of HCS and AEFS is that the character of the Mara Rosa deposit and the Mineral

Resource Estimate reported is robust and justifies engineering studies to investigate

economic feasibility.

25.2 Mining

The geotechnical work to date, and data provided, have been reviewed and summarised.

Coffey Mining’s review of these works suggests that they are sufficient for PFS level.

Groundwater and seismic effects, and operational considerations such as blasting, may affect

the PFS geotechnical slope design.

The mining study has demonstrated Proven and Probable Mineral Reserves of 17.17 Mt at a

grade of 1.72 g/t with contained gold totalling 945,200 oz. This study shows sufficient merit to

support implementation of feasibility study and mine development with a projected seven year

life of mine.

Proven Mineral Reserves comprise 31% of total reserves. Additional low grade resources

exist within the optimised pit and may be available for processing with an improved gold price.

Grade control will be essential during mining operations for reconciliation of production with

the block model and to maximise the cash flow. The evolution of gold price, operating costs

and plant recovery must be closely monitored to support ongoing management planning.

The estimated operating costs are compatible with similar operations in Brazil.



The design for the Tailings Storage Facility and related Water Storage Facility for the Mara

Rosa Project has been aimed at optimising tailings storage capacity by maximising tailings

density; and reducing environmental and societal impact.

Construction of the TSF will be in four stages over the life of the mine.

25.3 Metallurgy and Mineral Processing

Metallurgical testwork has identified a free milling gold (~75%) and refractory gold (~25%)

associated with sulphides and tellurides. Gold recoveries in excess of 93% can be achieved

with conventional carbon in leach (CIL) technology with the addition of pre-oxidation prior to

CIL to treat the refractory component.

With this metallurgical recovery, a total of 869,600 oz gold should be recovered during the

mine life described in this study.

The process flowsheet developed includes primary crushing followed by secondary and

tertiary crushing in closed circuit. Tertiary crushed material will feed mills with cyclone

classification to achieve the required P80 grind of 45 µm.

Tailings will be thickened to recover most of the cyanide in solution, with the remaining pulp

detoxified to remove free cyanide prior to disposal to the TSF.

25.4 Infrastructure

Plant site infrastructure includes a 138 kV power supply line and substation in the nearby city

of Porangatu.

Process plant water supply will integrate storage of old mine pit dewatering and surface runoff

in a water storage facility located between the plant site and the tailings disposal, together

with recovery from the TSF once production starts. Additional surface water sources are

available and are being investigated.

A preferred tailings storage facility site has been located south west of the plant site. The TSF

site is within 1 km of the plant site. A wet slurry (59% solids) tailings storage facility has been

designed for 20 Mt of tailings to be constructed in four stages over the life of mine followed by

vegetation coverage for final TSF closure. The TSF was designed based on preliminary

tailings storage requirements provided to Coffey Mining.

Auxilliary Mara Rosa Project buildings for administration, mine surface shops, and security

facilities will be constructed around the plant site. An operations campsite will be built close to

the plant site.



25.5 Capital and Operating Cost Estimates

Capital costs include direct and indirect project capital for the mine, process plant and

infrastructure. Project capital costs total US$189.5 M. The total indirect cost is US$32.6 M and

includes studies and management, rebate from power utility, insurance, contingency and

initial working capital. A 10% contingency is placed on initial direct and indirect capital costs

for the mine, plant and surface infrastructure. The total contingency allowance for the project

is US$16.9 M.

The Mara Rosa Project operating costs include fixed and variable costs for mine production,

plant production, tailings management and general and administrative services for the

operation.

Mine operating costs are estimated at US$12.59/t of ore, with a strip ratio of 8:1. However the

cost of waste rock pre-stripping in Year 0 has been capitalised for the purposes of economic

analysis. Consequently, mine operating costs average US$11.85 /t of ore during the

production Years 1 to 7 (excluding pre-stripping in Year 0 that are capitalised), with an

operational strip ratio of 7.4:1.

Plant operating costs, at a design processing rate of 2.5 Mtpa, total US$9.73 /t ore processed

including tailings disposal, and G&A costs average US$1.83 /t ore. These costs, pro-rated for

varying annual mine production rates, are US$9.78 /t and G&A costs of US$1.90 /t of ore

processed respectively.

Total cash operating costs are US$21.63 /t ore or US$464 /oz of gold produced, averaged

over life of mine Years 1 to 7.

25.6 Economic Analysis

A cash flow model incorporating Project and life of mine production, capital costs and

operating costs indicates that the Project has an after tax NPV of US$178.5 M at a discount

rate of 5%. A sensitivity analysis considering positive and negative variations of up to 30% in

either direction were applied independently to: gold price, operating cost and capital cost. The

results of the sensitivity analysis demonstrate that the project is most sensitive to variation in

gold price. Initial capital cost had the least impact on the sensitivity of the NPV.

25.7 Risk

25.7.1 Introduction

The risk analysis as described in this section is based on the Australian Standard for Risk

Management, AS 4360:2004 and an overview of the framework is provided in Figure

25.7.1_1.



Figure 25.7.1_1 Figure Description

AS/NZS 4360:2004

Mo

nit

or

/

Re

vie

w

Sta

ke

ho

lde

r C

on

sult

ati

on

/

Co

mm

un

ica

tio

n

Establish Goals & Context

Analyse Risks

Likelihood

Consequence

Evaluate the Risks

Treat the Risks

AS 4360:2004 states that risk analysis involves consideration of the sources of risk, their

consequences and the likelihood that those consequences may occur.

Qualitative measures of consequence and likelihood are described in Table 25.7.1_1 and

Table 25.7.1_2 respectively.

Based on the level of consequence and likelihood of an event as described above a risk

profile can be drawn up as per the matrix shown in Table 25.7.1_3.

Table 25.7.1_1

Mara Rosa Project

Qualitative Measures of Consequence (As per AS4360:2004)

Consequences Description

Catastrophic Very large financial loss; death or serious injury to multiple persons; major loss of plant resulting in >3 months loss of production capability; toxic environmental release off-site with detrimental effect.

Major Major financial loss; death or serious injury to multiple persons; extensive loss of plant resulting in 1-3 months loss of production capability; off-site environmental release without detrimental effect or on-site release with detrimental effect.

Moderate High financial loss; serious injury to multiple persons; moderate loss of plant resulting in 1



week to 1 month loss of production capability; on-site environmental release contained with assistance without causing long-term detrimental effect.

Minor Medium financial loss; minor injury to one or two persons; minor loss of plant resulting in 1 day to 1 week loss of production capability; on-site environmental release immediately contained without long-term detrimental effect.

Insignificant Low financial loss; no injuries; less than one-day loss of production capability; no environmental impact.

Table 25.7.1_2

Mara Rosa Project

Qualitative Measures of Likelihood (As per AS4360:2004)

Likelihood Description

Almost Certain Event is expected to occur in most circumstances; more than one event per month.

Likely Event will probably occur in most circumstances; less than one event per month but more than one event per year.

Possible Event might occur at some time; less than one event per year but more than one event per five years.

Unlikely Event could occur at some time; less than one event per five years.

Rare Event may only occur in exceptional circumstances or is unlikely to occur.



Table 25.7.1_3

Mara Rosa Project

Qualitative Risk Analysis Matrix (As per AS4360:2004)

Consequence

Likelihood 5 Catastrophic

4 Major

3 Moderate

2 Minor

1 Insignificant

Almost Certain Extreme Risk Extreme Risk High Risk Significant Risk Significant Risk

Likely Extreme Risk High Risk Significant Risk Significant Risk Moderate Risk

Possible High Risk High Risk Significant Risk Moderate Risk Low Risk

Unlikely High Risk Significant Risk Moderate Risk Low Risk Low Risk

Rare Significant Risk Significant Risk Moderate Risk Low Risk Low Risk

Legend Extreme risk - Immediate action required

High risk - Senior management attention needed

Significant risk - Management attention needed

Moderate risk - Management responsibility must be specified

Low risk - Manage by routine procedure

25.7.2 Risk Assessment

The risk assessment as described in this section pertains to mining aspects only and any

other, Project specific, risks have not been considered.

No event that was categorised as an Extreme Risk was identified. Events that have been

identified as High Risk are listed below:

Mining costs under-estimated;

Pit slope incorrect;.or

Reserve grade over-estimated.

A complete list of risks that have been assessed are summarised in Table 25.7.2_1.



Table 25.7.2_1

Mara Rosa Project

Summary of Risk Analysis

Risk Consequence Likelihood Level of Risk Comment / Mitigation Measure

Resource grade over-estimated by more than 15% Catastrophic Unlikely High Unlikely on a global scale, however, possible on a local scale.

Gold price over estimated by more than 15% Major Possible High Beyond the control of the operator

Rock density inadequately measured Minor Likely Significant Rock density measurement program in progress

Pit slope angle incorrect Catastrophic Possible High Geotechnical drill program for FS level study in progress

Reserve grade over-estimated by more than 15% Catastrophic Unlikely High Grade control program required

Reserve tonnes over-estimated by more than 15% Major Unlikely Significant Unlikely on a global scale, however, possible on a local scale.

Mining Capital costs underestimated Major Possible High Formal tender to be sought to firm up pricing

Plant Capital costs underestimated Major Possible High Formal tender to be sought to firm up pricing

Mining Capital costs underestimated Major Possible High Feasibility level study to determine

Plant Capital costs underestimated Major Possible High Feasibility level study to determine

Pit design unsuitable Major Unlikely Significant Main pit separated into north and south areas by saddle of Inferred Resources. Further resource drilling for FS

Production rate not achieved Moderate Unlikely Moderate Add equipment and or employ contractor.

Insufficient or incorrect equipment Moderate Rare Moderate If new equipment required then long lead times may result in lengthy Project delays. Employ contractor.

Project water supply inadequate Major Unlikely Significant Test work in progress.

TSF basin permeability unknown Moderate Possible Significant Test work to be undertaken.

Pit de-watering requirements under-estimated Moderate Possible Significant Test work to be undertaken.

Lack of available experienced mining personnel Moderate Possible Significant Recruitment and training priority.



26 RECOMMENDATIONS

A Feasibility Study is recommended for the Mara Rosa Project. A feasibility study work

program should include additional resource evaluation drilling, resource modelling and

estimation, geotechnical investigations, geotechnical site investigations for the plant, tailings

disposal and other key infrastructure, hydrogeology work to quantify and characterize mine

drainage and establish a water balance, metallurgical test work, detailed process design and

other standard components for a full feasibility study.

26.1 Feasibility Study Work Program

26.1.1 Geology

Additional exploration drilling is recommended to improve the confidence in inferred

resources, especially in that part of the deposit where the pit design separates the deposit into

North and South Pits.

Based on the new drilling, the Project’s resource model should be updated.

Samples for density determination should also be taken from mineralized zones. A target of at

least 30 to 40 density determinations per domain should be collected to adequately

characterize variability in the mineralized zone.

26.1.2 Pit Geotechnics

The requirements to achieve and the details of additional work required to meet Feasibility

Study level compliance for JORC and/or CIM Definition Standards reporting for the Mara

Rosa project are provided in Beer (2011).

Based on assessment of the geotechnical work done to date, it is Coffey Mining’s assessment

that a relatively small number of additional geotechnical bore holes would be required to

advance the geotechnical study to a FS level. The demonstrated geological continuity shown

by the sections provided by Amarillo to Coffey Mining suggests that drilling should be

concentrated in the southern part of the proposed pit, where it is at its deepest.

A total of six geotechnical bore holes have been recommended, including collar locations,

azimuths, declinations and design depths. A program of sample collection and

geomechanical testing has been proposed with sampling procedures to be confirmed during

an initial site visit.

In addition to the drilling and testwork outlined above, analytical and interpretative work will

also have to be undertaken on the data collected to meet the FS level requirements.

26.1.3 Mining Study




Whilst this report provides information on the proposed design, a significant portion of this

work is based on assumptions and these need to be verified before the design work

progresses to the next stage. Equally, there are aspects of the project design that have not

been addressed, where further detailed analysis is desirable. The details of investigations

and design work that should be carried out to progress the development of the TSF and WSF

are provided in Saunders (2011). These are summarised here as site investigation and

feasibility design.

The investigations that should be carried out prior to completion of the next stage of design

include:

Geotechnical site investigation for the TSF and WSF;

Drilling and equipping of the monitoring bores;

Drilling and permeability testing of investigation bores;

Laboratory testing of both tailings and waste rock for chemical content and static PAF;

Testing to determine the tailings geotechnical characterisation; and

Confirmation of geotechnical parameters for the embankment construction materials;

A factual report documenting these investigations should be compiled and appended to the

design report to substantiate the parameters used;

The following work should be undertaken to support the development’s application and

approval:

Modelling of the tailings beach;

Seepage and stability analyses using parameters from investigations;

Hydrological optimisation by conducting a site wide water balance;

Inflow and outflow hydrographs for spillways;.

Construction material balance;

Project implementation scheduling; and

Geotechnical risk assessment of the overall project including water and tailings management.

These investigations and aspects of the design will be incorporated into the feasibility design

report for the TSF and WSF, to provide a comprehensive document that will assist in

progressing the project.



26.1.5 Metallurgical Testwork

Further testwork is deemed necessary assuming the project proceeds to a feasibility level of

study. The testwork required at this stage would need to include variability testwork, as well

as confirmatory composite testwork and optimisation testwork on the grind versus pre-

oxidation characteristics of the mineralisation.

In addition to this, ancillary testwork requirements would need to be completed. This will

include confirmation that the site water contains no deleterious elements, settling/thickening

testwork and pulp viscosity testwork.

The details of the required testwork programme are included in Smith and Witt (2011).

26.1.6 Plant Design and Engineering

Results of the current confirmatory testwork may lead to fine adjustments to the process

parameters and possible variations in pre-oxidation and leach residence times in the final

plant design. There is also the possibility that fine grinding to a P80 of 45 μm may not be

necessary as classification in the plant will be by cyclone and implies that the heavier

sulphides/tellurides get preferentially recycled and reground. An attempt to quantify this is

being made in the confirmatory testwork programme.

With a better understanding of the water balance and in general water requirements, it is

hoped that a simpler more manageable water supply system can be introduced in the final

plant design. Although the local electricity distribution company has manifested that the

138 kV line from Porongatu would be the best method of supplying the plant, it is felt that

there could exist other cheaper alternatives and this will also be investigated during the

definitive study.

26.1.7 Environmental and Community

Continuing studies into environmental, hydrological, hydrogeological and community issues

will be required to fulfil Brazil legislation and regulation. The specific programs of work may

have significant lead times and early implementation is necessary. The environmental work

begun as part of this study is already progressing towards feasibility study level compliance

and regulatory approval.

26.2 Feasibility Study Program and Budget

The recommended work plan for the Feasibility Study began in July 2011 with the

implementation of the geotechnical and metallurgical testwork components. Hydrological and

environmental activities have also been ongoing throughout 2011. The main body of the

Feasibility Study work will commence soon after completion of this PFS, in December 2011.

The list of activities include:

Drilling (US$2.4M) to collect data and samples for:



resource model update

geotechnical characterization and condemnation of tailings and plant site locations

samples for metallurgy and tailings test work

Metallurgical testwork program (US$0.4M).

Geotechnical testwork program(US$0.4M)

Hydrogeological study (US$0.4M)

Tailings site testwork program (US$0.4M)

An updated Mineral Resource Model incorporating exploration data to improve

confidence in Mineral Resources (US$0.05M).

An updated mine design and mine schedule incorporating new hydrogeological, and

geotechnical data testwork. (US$0.5M).

Feasibility study including process and infrastructure design, engineering, capital and

operating cost estimation and financial analysis incorporating results of the geotechnical,

hydrogeological, mine design and mine schedule and metallurgical test work (US$1.5M)

Field expenses to continue with environmental base line study, property maintenance,

field staff and overheads (US$1.0M)

The recommended feasibility work plan will require a budget of approximately US$6.3M.



27 REFERENCES

Alho et al., 1986.

ANCOLD, 1999. ‘Guidelines on Tailings Dam Design, Construction and Operation’.

Beer, A. 2011. Geotechnical Report for Pre-feasibility Study. Mara Rosa Project. Prepared by

Coffey Mining Pty Ltd for Amarillo Mineracao do Brasil LTDA. pp 38.

BRASIL, 2005. Ministério do Meio Ambiente. Conselho Nacional do Meio Ambiente.

Resolução nº 357. Classificação dos corpos de água e diretrizes para o

enquadramento dos corpos de água superficiais e estabelecimento das condições e

padrões de lançamentos de efluentes. Diário Oficial da União, Brasília, DF, n. 53, 18

de março de 2005. Seção 1, p. 58-63.

Brazil, 2003. Ministério do Meio Ambiente/IBAMA (Ministry of the Environment/IBAMA)

Brazilian Society of Herpetology, 2010.

BVP Engenharia, 2011. Visita Tecnica e Avaliacao dos Dados Geotecnicos do Deposito

Posse – Projeto Mara Rosa. Report prepared for Amarillo Gold Corporation dated

February 2011.

Callisto, M., Moretti, M. & Goulart, M. ,2001. Macroinvertebrados bentônicos como

ferramenta para avaliar a saúde de riachos. Rev.Bras.Rec.Hídricos. 6: 71-82.


Economic Assessment, Mara Rosa Gold Property, Goias State, Brazil. Report

prepared for Amarillo Gold Corporation dated 29 February 2008.


Economic Assessment, Mara Rosa Gold Property, Goias State, Brazil. Report

prepared for Amarillo Gold Corporation dated 29 February 2008.

CBRO – Brazilian Committee of Ornithological Registers. 2010.

CIM Standing Committee on Reserve Definitions. 2010. CIM DEFINITION STANDARDS - For

Mineral Resources and Mineral Reserves. Canadian Institute of Mining, Metallurgy and

Petroleum. pp10.

Coffey, June 2011. ‘Future Tailings Storage Posse Deposit, Mara Rosa Project, Brazil’.

Colli et al., 2002.

Dias, B. F. S. Cerrados: uma caracterização. In: DIAS, B. F. S. (Coord.). Alternativa de

desenvolvimento dos cerrados: manejo e conservação dos recursos naturais

renováveis. Brasília, IBAMA, 1992, p. 11-25.

Fonseca, FL., Horta, A. 2011. Projeto da Mina Posse, Mining Study. 12 December 2011.



Coffey Mining.

Hidrovia. 2011. Estudos Hidrogeológicos e Hidrológicos Básicos ‘PFS’ Mina De Posse, Mara

/ Go. Prepared by HIDROVIA Hidrogeologia e Meio Ambiente Ltda. for Coffey Mining

Pty Ltd and Amarillo Gold Corporation. pp 59.

Hoogvliet Contract Services & Australian Exploration Field Services Pty Ltd. 2011. Report on

Independent Site Visit and Resource Estimate, Posse Deposit. Prepared for Amarillo

Gold Corporation. pp 162.

Hoogvliet Contract Services and Australian Exploration Field Services PL. 2011. Report on

Independent Site Visit and Resource Estimate. Posse Deposit, Mara Rosa, Goias

State, Brazil. Report prepared for Amarillo Gold Corporation dated 30 July 2011.

Hoogvliet Contract Services and Australian Exploration Field Services PL. 2010. Independent

Mineral Resource Estimate and Preliminary Economic Assessment, Posse Deposit,

Mara Rosa, Goias State, Brazil. Report prepared for Amarillo Gold Corporation dated

30 June 2010.

Hoogvliet Contract Services, 2010. Independent Mineral Resource Estimate and Preliminary

Economic Assessment, Posse Deposit, Mara Rosa, Goias State, Brazil. Report

prepared for Amarillo Gold Corporation dated 30 June 2010.

International Union for Conservation of Nature (IUCN), 2010.

Mittermeier, R. A. et al. 1998. Biodiversity hotspots and major tropical wilderness areas:

approaches to setting conservation priorities. Conservation Biology 12 (3): 516-520.

Myers, N. et al. 2000. Biodiversity hotspots for conservation priorities. Nature 403 (24): 853-

858.

Neotropica Tecnologia Ambienttal Ltda., 2010. Environmental Assessment of the Mara Rosa

Project Goiás State, Brazil, for Amarillo Mineraçao Do Brasil Ltda.

Neotropica Tecnologia Ambienttal Ltda., 2011. Environmental Assessment of the Mara Rosa

Project Goiás State, Brazil, for Amarillo Mineraçao Do Brasil Ltda.

Nogueira, M.G.; Matsumura-Tundisi, T. 1996. Limnologia de um sistema artificial raso

(Represa do Monjolinho - São Carlos, SP). Dinâmica das populações planctônicas.

Acta Limnologica Brasiliensia, Botucatu, v. 8, p. 149-168.

Onix Engenharia e Consultoria Ltda. 2011. Estudo de Pré-Viabilidade. Mara Rosa Project.

Prepared for and in collaboration with Amarillo Gold Corporation. pp 583.

Reynolds, C. S. et al. 2002. Towards a functional classification of the freshwater

phytoplankton. Journal of Plankton Research 24:417-428.

Rodrigues, 2005.



Round, 1993.

Saunders, C. 2011. PFS Report : Tailings Storage Facility and Water Storage Facility. Mara

Rosa : Posse Deposit. Prepared by Coffey Mining Pty Ltd for Amarillo Gold

Corporation. pp 47.

Silva & Bates, 2002.

Silva, 1995.

Smith, R., Witt, C. 2011. Mara Rosa Metallurgical Testwork Report. Mara Rosa Pre-

Feasibility Study. Prepared by Coffey Mining Pty Ltd for Amarillo Gold Corp. pp81.

William Freire. 2011. Synthetic analysis of the Exploration Permits referring the Mara Rosa

Project. Letter opinion on the legal status of mineral tenure by William Freire,

Advocados Associados, Belo Horizonte, MG, Brazil.

mara rosa project, goiás state, brazil (latitude …av. afonso pena, 4001 12º andar - bairro serra...

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