NI 43-101 Technical Report Updated Feasibility Study Aurora Gold Project

8/19/2019 NI 43-101 Technical Report Updated Feasibility Study Aurora Gold Project

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NI 43-101 Technical Report

Updated Feasibility Study

Aurora Gold Project

Guyana, South America



NI 43-101 Technical ReportUpdated Feasibility Study Aurora Gold Project Guyana, South America

C O N T E N T S

1.0 SUMMARY .......................................................................................................1

INTRODUCTION .......................................................................................................... 11.1

PROPERTY DESCRIPTION .............................................................................................. 11.2

GEOLOGY AND MINERALIZATION ................................................................................... 21.3 MINERAL RESOURCE ESTIMATE ..................................................................................... 21.4

OPEN PIT MINING ...................................................................................................... 31.5

UNDERGROUND MINING ............................................................................................. 61.6

COMBINED MINERAL RESERVE ESTIMATE AND PRODUCTION SCHEDULE .............................. 81.7

METALLURGY ............................................................................................................. 91.8

MINERAL PROCESSING ................................................................................................ 91.9

PROJECT INFRASTRUCTURE ......................................................................................... 101.10

SOCIAL AND ENVIRONMENTAL ASPECTS ....................................................................... 111.11

CAPITAL AND COST ESTIMATES ................................................................................... 141.12

FINANCIAL ANALYSIS ................................................................................................. 151.13

CONCLUSION AND RECOMMENDATIONS ....................................................................... 161.14

2.0 INTRODUCTION.............................................................................................. 21

BACKGROUND INFORMATION 212 1




PHYSIOGRAPHY ........................................................................................................ 295.4

6.0 HISTORY ........................................................................................................ 32

EXPLORATION WORK PRIOR TO GUYANA GOLDFIELDS .................................................... 326.1

EXPLORATION BY GUYANA GOLDFIELDS (1998 TO 2009) ............................................... 346.2

PREVIOUS MINERAL RESOURCE ESTIMATES ................................................................... 356.3

7.0 GEOLOGICAL SETTING AND MINERALIZATION ................................................ 38

REGIONAL GEOLOGY ................................................................................................. 387.1

PROPERTY GEOLOGY ................................................................................................. 397.2

MINERALIZATION ...................................................................................................... 417.3

7.3.1 RORY’S KNOLL .......................................................................................... 42

7.3.2 ALECK HILL .............................................................................................. 44

7.3.3 WALCOTT HILL ......................................................................................... 47

7.3.4 MAD KISS ................................................................................................ 47

8.0 DEPOSIT TYPES............................................................................................... 50

9.0 EXPLORATION ................................................................................................ 51

10.0 DRILLING........................................................................................................ 52

DOWNHOLE SURVEYING ............................................................................................ 5510.1

DRILLING PATTERN AND DENSITY ................................................................................ 5510.2

DRILL CORE SAMPLING .............................................................................................. 5610.3

11.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY ........................................ 57

SAMPLE PREPARATION AND ANALYSES ......................................................................... 5711.1

11 1 1 SAMPLE PREPARATION AND ANALYSES PRIOR TO 2009 57




13.2.1 SAMPLE SELECTION ................................................................................... 66

13.2.2 HEAD ASSAY ............................................................................................ 67

13.2.3 SAMPLE PREPARATION ............................................................................... 68TESTWORK REVIEW ................................................................................................... 6913.3

13.3.1 MINERALOGY ........................................................................................... 69

13.3.2 COMMINUTION CHARACTERISTICS ............................................................... 70

13.3.3 GRAVITY CONCENTRATION ......................................................................... 72

13.3.4 LEACHING ................................................................................................ 74

13.3.5 RHEOLOGY AND SETTLING THICKENING ......................................................... 79

13.3.6 CYANIDE DETOXIFICATION .......................................................................... 81

13.3.7 METALLURGICAL RECOVERIES ..................................................................... 8213.3.8 CONCLUSIONS .......................................................................................... 82

14.0 MINERAL RESOURCE ESTIMATES .................................................................... 83

MINERAL RESOURCE ESTIMATION METHODOLOGY ........................................................ 8314.1

DATABASE ............................................................................................................... 8314.2

14.2.1 GENERAL ................................................................................................. 83

14.2.2 DATA VALIDATION .................................................................................... 83

RESOURCE MODELING PROCEDURES ............................................................................ 8414.3

14.3.1 GEOLOGICAL MODEL ................................................................................. 84

14.3.2 DATABASE PREPARATION ........................................................................... 86

14.3.3 PREPARATION OF ASSAY COMPOSITES, OUTLIER ANALYSES AND STATISTICS ....... 86

14.3.4 SPECIFIC GRAVITY ..................................................................................... 93

14.3.5 VARIOGRAPHY .......................................................................................... 93

14.3.6 BLOCK MODEL ......................................................................................... 95

14.3.7 GRADE INTERPOLATION ............................................................................. 95

14 3 8 RESOURCE MODEL VALIDATION 98




16.2.9 OPEN PIT PRODUCTION SCHEDULE ............................................................ 146

16.2.10 OPEN PIT OPERATION .............................................................................. 149

16.2.11 OPEN PIT EQUIPMENT FLEET .................................................................... 15316.2.12 OPEN PIT WORKFORCE ............................................................................ 154

UNDERGROUND MINING ......................................................................................... 15516.3

16.3.1 INTRODUCTION ....................................................................................... 155

16.3.2 RESOURCE MODEL .................................................................................. 155

16.3.3 UNDERGROUND MINE GEOTECHNICAL ....................................................... 156

16.3.4 NUMERICAL MODELING RESULTS .............................................................. 171

16.3.5 MINING METHOD ................................................................................... 179

16.3.6 MINE DESIGN ......................................................................................... 18016.3.7 CUT-OFF GRADE ..................................................................................... 183

16.3.8 MINE DEVELOPMENT .............................................................................. 186

16.3.9 MINE PRODUCTION ................................................................................. 190

16.3.10 UNDERGROUND DEVELOPMENT AND PRODUCTION SCHEDULE ....................... 192

16.3.11 UNDERGROUND EQUIPMENT FLEET ........................................................... 194

16.3.12 UNDERGROUND PERSONNEL ..................................................................... 194

16.3.13 MATERIAL HANDLING .............................................................................. 197

16.3.14 MINE SAFETY ......................................................................................... 19716.3.15 MINE SERVICES ...................................................................................... 198

16.3.16 VENTILATION ......................................................................................... 202

16.3.17 AIR COOLING ......................................................................................... 209

16.3.18 DEWATERING ......................................................................................... 213

16.3.19 COMMUNICATIONS ................................................................................. 217

16.3.20 EXPLOSIVES STORAGE AND INITIATION SYSTEM ............................................ 218

16.3.21 FUEL STORAGE AND DISTRIBUTION ............................................................ 218

16 3 22 COMPRESSED AIR 218




17.6.7 TAILINGS MANAGEMENT AREA (TMA) ...................................................... 228

17.6.8 RAW AND PROCESS WATER ...................................................................... 228

17.6.9 REAGENTS ............................................................................................. 229PROCESS CONTROL PHILOSOPHY ............................................................................... 23117.7

PROCESS FACILITY INFRASTRUCTURE AND SERVICES ...................................................... 23217.8

17.8.1 AIR ....................................................................................................... 232

17.8.2 WATER ................................................................................................. 232

PROCESS FACILITY ANCILLARY BUILDINGS ................................................................... 23317.9

EQUIPMENT SIZING ................................................................................................. 23317.10

17.10.1 PRIMARY CRUSHING ................................................................................ 23317.10.2 SECONDARY AND TERTIARY CRUSHING ....................................................... 233

17.10.3 TERTIARY CRUSHING ................................................................................ 233

17.10.4 GRINDING CIRCUIT .................................................................................. 234

17.10.5 PRELEACH THICKENING AND CIL ................................................................ 234

17.10.6 CARBON ELUTION ................................................................................... 234

17.10.7 CYANIDE DETOXIFICATION ........................................................................ 235

18.0 PROJECT INFRASTRUCTURE .......................................................................... 236

PROJECT LOGISTICS ................................................................................................. 23618.1

ON-SITE INFRASTRUCTURE ....................................................................................... 23618.2

18.2.1 POWER PLANT AND DISTRIBUTION ............................................................. 236

18.2.2 ON SITE ROADS ...................................................................................... 237

18.2.3 UTILITIES AND SERVICES ........................................................................... 238

18.2.4 SITE BUILDINGS AND FACILITIES ................................................................. 239

18.2.5 TAILINGS MANAGEMENT AREA ................................................................. 241

18 2 6 FRESH WATER POND 242




20.3.4 WATER MANAGEMENT ............................................................................ 264

PERMITTING .......................................................................................................... 26520.4

SOCIAL OR COMMUNITY RELATIONS REQUIREMENTS .................................................... 26620.5

MINE CLOSURE ...................................................................................................... 26820.6

20.6.1 GENERAL DESCRIPTION OF MINE RECLAMATION AND CLOSURE PLAN .............. 268

20.6.2 SUMMARY OF SITE CLOSURE AND WASTE DISPOSAL STRATEGY ...................... 269

20.6.3 COST ESTIMATE ...................................................................................... 271

20.6.4 POST-CLOSURE MONITORING ................................................................... 272

21.0 CAPITAL AND OPERATING COSTS ................................................................. 274

CAPITAL COSTS ...................................................................................................... 27421.121.1.1 OPEN PIT MINE ...................................................................................... 274

21.1.2 UNDERGROUND MINE ............................................................................. 275

21.1.3 PROCESS, ON- AND OFF-SITE INFRASTRUCTURE ........................................... 276

21.1.4 OWNER COSTS & MINE CLOSURE .............................................................. 277

OPERATING COSTS .................................................................................................. 27821.2

21.2.1 OPEN PIT MINE ...................................................................................... 278

21.2.2 UNDERGROUND MINE ............................................................................. 279

21.2.3 CIL PROCESS PLANT ................................................................................ 280

21.2.4 GENERAL & ADMINISTRATIVE COSTS .......................................................... 282

22.0 ECONOMIC ANALYSIS ................................................................................... 284

PRINCIPAL ASSUMPTIONS ........................................................................................ 28422.1

22.1.1 TECHNICAL PARAMETERS ......................................................................... 284

22.1.2 PRODUCTION SUMMARY .......................................................................... 285

CASH FLOW ........................................................................................................... 28622.2






Table 11–2: Specific Gravity Database for the Aurora Gold Project ......................................... 60 Table 12–1: Summary of Analytical Quality Control Data Produced by Guyana

Goldfields on the Aurora Gold Project between November 30, 2010 andJuly 31, 2011 .......................................................................................................... 63 Table 13–1: List of Metallurgical Reports .................................................................................. 65 Table 13–2: List of SRK Litho-tectonic Domains in Relation to Guyana Goldfields Logs ........ 66 Table 13–3: Boreholes Co-ordinates .......................................................................................... 66 Table 13–4: Head Analysis .......................................................................................................... 68 Table 13–5: Lithology and Mineralization by Ore Zone, Aurora Gold Project (Mineral

Resource Evaluation, SRK) ..................................................................................... 69 Table 13–6: Crystalline Mineral Assemblage Phases of the Saprolite and Fresh Rock

Composite Samples (12088-001) ......................................................................... 70 Table 13–7: Bond Ball Mill Grindability Test Summary (12088-001) ...................................... 70 Table 13–8: Comminution Test Summary (12088-002) ........................................................... 70 Table 13–9: Comminution Test Summary (12088-005) ........................................................... 71 Table 13–10: Mill Model Grinding Media Wear Prediction10 ....................................................... 72 Table 13–11: Gravity Separation Test Results (12088-001) ...................................................... 72 Table 13–12: Gravity Separation Test Result (12088-002) ........................................................ 72 Table 13–13: Gravity Separation Test Results (12088-005) ...................................................... 73 Table 13–14: Gravity Separation Test Results (12088-005 - Knelson only) ............................. 73

Table 13–15: Intensive Cyanidation Results Summary ............................................................... 74 Table 13–16: Whole Ore Cyanidation Results Summary (12088-001) ...................................... 74 Table 13–17: Bulk Whole Ore Cyanidation Test Results (12088-005) ...................................... 75 Table 13–18: Summary of Cyanidation Test Results on the Saprolite and Fresh Rock

Composites Gravity Tailings (12088-001) ............................................................ 75 Table 13–19: Gravity Tailings Cyanidation Results (12088-002) ............................................... 76 Table 13–20: Cyanidation of Gravity Tailings Test Results – Effect of Grind Size and

Lead Nitrate Addition (12088-005) ....................................................................... 76

T bl 13 21 C id ti f G it T ili g T t R lt Eff t N CN C t ti




Table 14–12: Impact of the 2012 Drilling Program on the Resource Statement .................... 105 Table 14–13: Global Block Model Quantity and Grades* Estimates at Various Cut-off

Grades, Aurora Gold Project, Guyana .................................................................. 106 Table 15–1: Mineral Reserve Statement*, Aurora Gold Project, Guyana, SRK

Consulting (Canada) Inc., January 11, 2013 ...................................................... 107 Table 16–1: Hydraulic Parameters of Geologic Units in Groundwater Flow Model ............... 110 Table 16–2: Open Pits Excavation Schedule ........................................................................... 113 Table 16–3: Summary for the Material Testing of Each Rock Type (AMEC, 2012). .............. 121 Table 16–4: Summary of Major Joint Sets per Domain of Rory’s Knoll (AMEC, 2012). ........ 122 Table 16–5: Statistical Analysis of Rock Mass Properties from Core Logging (AMEC,

2012) ..................................................................................................................... 123

Table 16–6: Summary of the High Wall SLIDE Analyses for Segment 1, 4 and 5 (AMEC,2012). .................................................................................................................... 126

Table 16–7: Recommended Slope Geometry Based on Kinematic and Slope StabilityAnalyses (AMEC, 2012) ........................................................................................ 126

Table 16–8: Recommended Slope Geometry .......................................................................... 131 Table 16–9: Aurora Pit Optimization Input Parameters ........................................................... 133 Table 16–10: Overall Slope Angles used in the Pit Optimization .............................................. 134 Table 16–11: Whittle Pit Shells ................................................................................................... 137 Table 16–12: Pit Design Parameters .......................................................................................... 138

Table 16–13: External Dilution, Dilution Grades and Mining Losses for Each MineralizedZone ....................................................................................................................... 143

Table 16–14: Economic Parameters for CoG Estimate ............................................................. 143 Table 16–15: Open Pit Reserves ................................................................................................. 144 Table 16–16: Proven and Probable Reserves ............................................................................ 144 Table 16–17: Open Pit Mine, Ore and Waste By Pit Phase ....................................................... 145 Table 16–18: Open Pit Production Contributions by Area ......................................................... 146 Table 16–19: Open Pit Production Schedule ............................................................................. 147

T bl 16 20 O Pit Mi i g E i t A il bilit 150




Table 20–1: Potential Environmental and Social Impacts and AssociatedManagement/Mitigation Strategies ..................................................................... 250

Table 20–2: Summary of Major Permit and License Requirements ....................................... 265 Table 20–3: Estimated End of Mine Life Closure Costs .......................................................... 271 Table 21–1: LoM Project Capital ($000s) ................................................................................ 274 Table 21–2: Open Pit Capital ($000s)1 .................................................................................... 275 Table 21–3: Underground Capital (US$000s) .......................................................................... 276 Table 21–4: Process, On- & Off-Site Infrastructure Capital ($000s) ...................................... 277 Table 21–5: Owner & Closure Capital ($000s) ........................................................................ 278 Table 21–6: LoM Operating Costs ............................................................................................ 278 Table 21–7: Open Pit Operating Costs (LoM) ........................................................................... 279

Table 21–8: Underground Operating Costs (LoM) ................................................................... 280 Table 21–9: CIL Plant Operating Costs (LoM) .......................................................................... 280 Table 21–10: CIL Plant Power Requirements ............................................................................. 281 Table 21–11: Reagent Costs ....................................................................................................... 281 Table 21–12: Consumable Costs ................................................................................................ 282 Table 21–13: General & Administrative Costs ........................................................................... 283 Table 22–1: TEM Principal Assumptions .................................................................................. 285 Table 22–2: Production Summary ............................................................................................ 286 Table 22–3: Technical-Economic Results ................................................................................. 287

F I G U R E S

Figure 4–1: Property Location Map ............................................................................................ 24 Fi 4 2 A G ld P j t L d T M LIDAR d SRTM D t 26




Figure 14–3: Sample Length Histograms for Aleck Hill North, Rory’s Knoll East, Walcott

and Walcott East. .................................................................................................... 88 Figure 14–4: Sample Length Histograms for Mad Kiss, Mad Kiss West and Mad Kiss

South. ....................................................................................................................... 89 Figure 14–5: Cumulative Frequency Plots for Gold Composites in Aleck Hill, Aleck Hill

High Grade, Rory’s Knoll and Rory’s Knoll High Grade - Capping Grades asIndicated .................................................................................................................. 90

Figure 14–6: Cumulative Frequency Plots for Gold Composites in Aleck Hill North,Rory’s Knoll East, Rory’s Knoll High Grade, Walcott and East Walcott -Capping Levels as Indicated. ................................................................................. 91

Figure 14–7: Cumulative Frequency Plots for Gold Composites in Mad Kiss, Mad Kiss

South and Mad Kiss West - Capping Levels as Indicated. ................................... 92 Figure 14–8: Aurora Gold Project Modeled Domains in Relation to the Conceptual Pit

Shells ..................................................................................................................... 101 Figure 14–9: Aurora 10 m Saprolite mineralization in Relation to the Conceptual Pit

Shell. ...................................................................................................................... 102 Figure 14–10: Aurora Gold Project Grade-Tonnage Curves. ....................................................... 106 Figure 16–1: Base Map for Guyana Site .................................................................................... 109 Figure 16–2: Measured Kh and Modeled Kx in Unconsolidated Deposits Based on

Distance from River .............................................................................................. 111

Figure 16–3: Measured Kh and Modeled Kx in Weathered Bedrock Based on Distancefrom River .............................................................................................................. 111

Figure 16–4: Geometric Mean of Kh and Modeled Kx vs. Depth ............................................. 112 Figure 16–5: Simulated Open Pits and Underground Mine Workings ..................................... 112 Figure 16–6: Base Case Scenario: Simulated Inflow Rates to Various Open Pits .................. 114 Figure 16–7: Base Case Scenario: Simulated Inflow Rates to SLR Workings ......................... 115 Figure 16–8: Sensitivity Analysis: Inflow to Rory's Knoll Pit ...................................................... 115 Figure 16–9: Sensitivity Analysis: Inflow to Aleck Hill Pit .......................................................... 116

Fi 16 10 S iti it A l i I fl t SLR W ki 116




Figure 16–31: Fracture frequency (fractures per metre) data ranges for thegeomechanical domains. ..................................................................................... 158

Figure 16–32: Location of boreholes for 2012 Drilling Investigation, geomechanical(above) and hydrogeological (below). .................................................................. 160

Figure 16–33: Updated geological model with underground geomechanical units.................. 162 Figure 16–34: Typical examples of Tonalite Domain. ................................................................. 163 Figure 16–35: Typical examples of Sericite Schist (SCS) Domain. ............................................ 164 Figure 16–36: Typical examples of Mafics Domain. ................................................................... 165 Figure 16–37: Typical examples of Interbedded Domain. .......................................................... 166 Figure 16–38: UCS strengths by Geomechanical Domain .......................................................... 167 Figure 16–39: Example of incidence of foliation angle taken from sample photos. ................ 167

Figure 16–40: Distinct strength groupings of the UCS test results. ........................................... 168 Figure 16–41: Isometric view - FLAC3D mining stages modeled. .............................................. 169 Figure 16–42: Mechanical and hydrological model interaction. ................................................ 170 Figure 16–43: Modeled mining induced vertical displacement (at ground surface and

50m below surface) during mid-life (above) and end-life (below) miningstages. Note: blue region represents vertical displacement between 1 and1.5 cm. ................................................................................................................... 171

Figure 16–44: Modeled mining induced relaxation at an end-life mining stage at 100mintervals below ground surface. ........................................................................... 172

Figure 16–45: Plan view - Modeled pore water pressure distribution around the mine atan end-life mining stage at -400m below ground surface. ................................ 173

Figure 16–46: Modeled stress and displacement encountered at mine infrastructure atan end-life mining stage at -900m below ground surface. ................................ 174

Figure 16–47: Empirical ground support design (Grimstad and Barton, 1993) ....................... 176 Figure 16–48: Rory’s Knoll Mine Design Looking South ............................................................. 181 Figure 16–49: Rory’s Knoll Mine Design Looking North ............................................................. 182 Figure 16–50: Typical sublevel plan view - zone of disturbance ................................................ 185

Fi 16 51 St dil ti ith i i i i i d th 186




A B B R E V I A T I O N S A N D A C R O N Y M S

atomic absorption spectrometry ........................................................................... AA

above mean sea level ............................................................................................. amsl

acre ......................................................................................................................... ac

ampere ................................................................................................................... A

annum (year) .......................................................................................................... a

billion ...................................................................................................................... Bbillion tonnes .......................................................................................................... Bt

billion years ago...................................................................................................... Ga

Bond Ball mill Work Index ...................................................................................... BWI

Bond Rod mill work Index ...................................................................................... RWI

British thermal unit ................................................................................................ BTU

Canadian Institute of Mining, Metallurgical, and Petroleum ................................ CIMCarbon-in-column ................................................................................................... CIC

Carbon-in-pulp ....................................................................................................... CIP

Carbon-in-leach ...................................................................................................... CIL

Closed circuit television .......................................................................................... CCTV

centimetre .............................................................................................................. cm

Crusher Work Index ................................................................................................ CWI






kilopascal ................................................................................................................ kPa

kilotonne................................................................................................................. kt

kilovolt(s) ................................................................................................................ kV

kilovolt-ampere ...................................................................................................... kVA

kilowatt ................................................................................................................... kW

kilowatt hour .......................................................................................................... kWh

kilowatt hours per tonne ........................................................................................ kWh/t

kilowatt hours per year .......................................................................................... kWh/a

less than.................................................................................................................. <life of mine.............................................................................................................. LoM

litre (liter) ............................................................................................................... L

litres per minute ..................................................................................................... L/m

megabytes per second ........................................................................................... Mb/s

megapascal ............................................................................................................. MPa

megavolt-ampere ................................................................................................... MVAmegawatt ............................................................................................................... MW

metre (meter) ......................................................................................................... m

metres above sea level .......................................................................................... masl

metres Baltic sea level ............................................................................................ mbsl

metres per minute.................................................................................................. m/min

metres per second .................................................................................................. m/s




pascal ...................................................................................................................... Pa

centipoise ............................................................................................................... mPa∙s

parts per million ..................................................................................................... ppm

parts per billion ...................................................................................................... ppb

percent ................................................................................................................... %

Protocol Independent Multicast ............................................................................ PIMS

pound(s) ................................................................................................................. lb

pounds per square inch .......................................................................................... psi

Probable maximum flood ....................................................................................... PMFreverse circulation drilling method ........................................................................ RC

revolutions per minute ........................................................................................... rpm

run of mill ............................................................................................................... RoM

semi-autogenous grinding ...................................................................................... SAG

second (plane angle) .............................................................................................. "

second (time) .......................................................................................................... sSGS Mineral Services .............................................................................................. SGS

short ton (2,000 lb) ................................................................................................. st

short tons per day .................................................................................................. st/d

short tons per year ................................................................................................. st/y

specific gravity ........................................................................................................ SG

square centimetre .................................................................................................. cm2

2




week ....................................................................................................................... wk

weight/weight ........................................................................................................ w/w

wet metric ton ........................................................................................................ wmt




U N I T S O F M E A S U R E

All dollars are presented in US dollars unless otherwise noted. For the purpose of this report the

exchange rates are US$1.00 = CDN$1.00 expect as needed for the sensitivity analysis. Common

units of measure and conversion factors used in this report include:

Linear Measure

1 inch = 2.54 centimeters

1 foot = 0.3048 meter

1 yard = 0.9144 meter

1 mile = 1.6 kilometers

Area Measure

1 acre = 0.4047 hectare

1 square mile = 640 acres = 259 hectares

Capacity Measure (liquid)

1 US gallon = 4 quarts = 3.785 liter

1 cubic meter per hour = 4.403 US gpm



NI 43-101 Technical ReportUpdated Feasibility Study | Aurora Gold Project |Guyana, South America

1 . 0 S U M M A R Y

I N T R O D U C T I O N 1.1

The Aurora Gold Project (or the “Project”) is a development stage gold project situated in the forested

region of Guyana, South America. Guyana Goldfields Inc. (“Guyana Goldfields” or the “Company””) holds

a 100% interest in the Project.

Guyana Goldfields has conducted extensive surface drilling, outlining several gold deposits in the Aurora

Gold Project area that are amenable to open pit and/or underground mining methods. Several resource

estimates have been completed with the most recent results announced publicly by Guyana Goldfields

on June 25, 2012.

Prior to the current study, SRK Consulting (Canada) Inc. (“SRK”) was commissioned by Guyana Goldfields

to prepare a Feasibility Study for the Aurora Gold Project. The NI 43-101 Technical Report, Feasibility

Study was issued by SRK on April 9, 2012.

In May 2012, Tetra Tech, Inc. (Tetra Tech) was commissioned by Guyana Goldfields to prepare this

Updated Feasibility Study for the Aurora Gold Project. For the purposes of the study, SRK prepared open

pit mine design, underground mine design and resource evaluation. Itasca International Inc. provided

groundwater hydrology input to SRK; Bluhm Burton Engineering provided the mine ventilation and

cooling input to SRK; Tetra Tech prepared the process design and site infrastructure and Technical-

Economic Model; and Environ International provided environmental evaluations.




The A1 prospecting licence (granted Prospecting Licence No. 14/2004) that encompasses the Aurora

Gold Project was replaced by a Mining Licence in November 2011. The licence gives Guyana Goldfields

the right to build and operate the mine. When the licence was issued, the Company also signed aMineral Agreement with the Government of Guyana and the Guyana Geology and Mines Commission

which sets the fiscal regime, taxation and royalties as they affect the operation of the mine. The licence

and mineral agreement were signed by Guyana Goldfields and the Company’s wholly owned subsidi ary

in Guyana, AGM Inc., and are valid for 20 years and renewable on application for an additional 7 year

period. The boundaries of the license form an oblong shape trending approximately southeast from the

south bank of the Cuyuni River. The Company received its Environmental Permit in from the Guyana

Environmental Protection Agency on September 27, 2010.

G E O L O G Y A N D M I N E R A L I Z A T I O N 1.3

The mineral resources reported herein are confined within an approximately 2 km long corridor, known

as the “Golden Square Mile” within the A1 licence. The Golden Square Mile area of the Aurora Gold

Project comprises folded metasedimentary and metavolcanic rock of the lower Cuyuni Formation that

has been metamorphosed to greenschist assemblages. The Golden Square Mile is located within a broad

regional, northwest trending, high strain zone characterized by strong northwest trending and sub-

vertical foliation and dip slip shearing (southwest over northeast) and strain partitioning intointerconnected network of discrete shear zones.

Gold mineralization at the Aurora Gold Project exhibits features analogous to mesothermal or

“orogenic” gold deposits in the West-African Palaeoproterozoic Birimian Supergroup, with all gold

mineralization controlled by a series of northwest trending shear zones.

The Aurora Project area is divided into four major areas of gold mineralization; Rory’s Knoll (includes




#1416). Ms. El-Rassi, Mr. Chartier, and Mr. Cole are independent Qualified Persons as this term is

defined in National Instrument 43-101. The effective date of this resource estimate is June 25, 2012.

Mineral resources are reported at two cut-off grades to reflect the fact that parts of the gold

mineralization are amenable for open pit extraction, while other parts are more likely amenable for

underground extraction. The consolidated Mineral Resource Statement for the Aurora Gold Project is

presented in Table 1 –1.

Table 1–1: Consolidated Mineral Resource Statement* Aurora Project,

Guyana, SRK Consulting (Canada) Inc., June 25, 2012

Classification Quantity Grade Contained Au 000’ Tonnes Au g/t 000’ Ounces Open Pit Mining Measured 5.77 3.23 0.60Indicated 27.01 2.49 2.16Inferred 5.12 1.54 0.25Underground MiningMeasured 0.00 0.00 0.00Indicated 30.06 3.91 3.78Inferred 11.81 4.12 1.56Combined MiningMeasured 5.77 3.23 0.60Indicated 57.06 3.24 5.94Inferred 16.93 3.34 1.82

* Mineral resources are not mineral reserves and do not have demonstratedeconomic viability. All figures have been rounded to reflect the relative accuracy ofthe estimates. The cut-off grades are based on a gold price of US$1,300 perounce of gold and metallurgical recoveries of ninety-five percent for saprolite andfresh material. Open pit resources are reported at a cut-off grade of 0.30 and 0.40gpt gold within conceptual pit shells for saprolite and fresh rock respectively,




inter-ramp slope angles of 51º to 56º has been defined for individual sectors of the pits with bench face

angles varying from 70˚ to 75˚, catch bench widths varying from 6 m to 10m, and final wall bench

heights of 15m. Overall pit slopes depending on the number and width of ramp sections vary from 40˚to 53˚.

The overburden soils are made of typically 5 m to 40 m thick saprolite and saprolitic rock (saprock). A

bench face angle (BFA) based on site experience of 70° is recommended, which was found to be stable

at site for excavations 5m to 7m high. The recommended inter-ramp angles vary from 30˚ to 36˚.

Open pit mining of near surface saprolite and fresh bedrock mineralization is planned. At Rory’s Knoll

the mineralization occurs within a distinctive tonalite zone that is more than 100m in diameter in the

open pit area of interest. In the other deposits, the mineralization occurs within a vein stockwork that

typically strikes southeast-northwest with a dip of about 75° to the southwest. Individual veins range

from 2m to about 20m in thickness.

Pit Optimization

Pit optimization was conducted using Whittle™ software based on a gold price of US$1,300/oz to create

a series of nested pit shells for analysis. The optimization in the Rory’s Knoll pit area was conducted as

an open pit/underground cross over which accounts for the opportunity to mine ore blocks by anunderground mining method. The remaining mineralized zones were evaluated with the conventional

Lerch-Grossman algorithm. The incremental tonnes of ore and waste, gold grade and net present value

in each sequential nested pit shell were analyzed in order to select the ultimate pit shell.

Mine design is based on a conventional surface mine operation using 152 mm blast holes, 7.7 m 3 front

end loaders for ore and waste loading; and haulage by a fleet of 43.5 tonne capacity trucks.




Upper Saprolite – 17%

Saprolite Veins - 23%

Total mineral reserves within the five designed pits are 13.7 Mt at an average grade of 2.55 g/t Au.

Waste rock mining totals 63.7 Mt for an overall strip ratio of 4.7 to 1.

Pre-production is planned to begin in late 2014 with production beginning in early 2015. The Rory’s Knoll

pit will be mined first in order to exploit the low strip ratio and to make the pit available to underground

mining as early as possible.

Open pit ore will be mined at a nominal rate of 1.75 Mt per year from 2015 to 2023. The designed

waste rock stockpiles are located as close as possible to each mining area in order to maximize haul

truck productivity. There is a total of 63.7 Mt of fresh and saprolite waste material mined.

Grade Control and Dewatering

During fresh rock mining, it is expected that gold mineralization will be recognizable in the mining face

by pit geologists and grade control technicians; although, grade control sampling will be necessary to

identify the zones above the cut-off grade. The upper saprolite mineralized zones are not expected to be

recognizable in the mining face due to extensive weathering, and identification of the location of zonesabove the cut-off grade will rely primarily on grade control definition drilling and sampling.

Rainfall that falls in the open pits plus pit groundwater inflow will be pumped from the pits to surface

sumps strategically located around the ultimate pit rim. An overland pumping system will transport the

water from these three surface sumps to the mine water pond located southeast of the Aleck Hill pit.

Manpower




U N D E R G R O U N D M I N I N G 1.6

Geomechanical

Detailed geomechanical site investigation work to a feasibility study level has been completed for the

underground mine at Rory’s Knoll and geomechanical design criteria as well as numerical modeling

inputs were developed. SRK completed a field program and updated the geotechnical domains in 2012

for the selected sublevel retreat (SLR) mining method.

Numerical modeling was performed by SRK and Itasca Denver using 3D finite element codes to create a

coupled geomechanical and groundwater flow model for open pit and underground mining of the Rory’s

Knoll deposit. The modeling objectives were to model stress, mining induced relaxation and disturbanceof the host rock mass, provide ground support recommendations, underground infrastructure stability

analysis, and water inflow estimates into the mine openings.

The Rory’s Knoll deposit is contained within the tonalite pipe and measures approximately 140 m x 100

m in plan and has been defined from surface to approximately 1,600 m below surface. The tonalite pipe

plunges to the north-west at around 80° and is the predominant gold bearing orebody. Two strongly

foliated north-west, south-east striking sericite shears are located on either side of the tonalite pipe.

Gold mineralization occurs along the shear contacts. The majority of the capital development will belocated in the Interbedded Geotechnical domain which is located to the north of the sericite shear and

south of the Cuyuni River. The rock quality of the tonalite pipe is good and has a rock mass rating

(RMR90) between 70 to 75 and an average uniaxial compressive strength (UCS) of 150 MPa. The rock

quality of the sericite shear is fair and has an RMR90 between 55 to 60 and an average UCS of 105 MPa.

The rock quality of the Interbedded domain is good and has an RMR 90 rating between 65 to 70 and an

average UCS of 110 MPa.




The mine design is based on decline access with rubber-tired diesel powered equipment. Development

and production will be achieved with a fleet of 55 tonne capacity haul trucks, 17 tonne loaders, jumbo

drills, production drills, mechanized bolters, ANFO and emulsion blasting machines and various supportequipment.

The mine has been designed with sublevels spaced at 25 m vertically apart with 15 m wide stopes.

Stopes will have a length approximately the width of the orebody ranging between 15 m to 120 m and

each sublevel will have between 8 and 11 stopes. Three sublevels are planned to be in production at a

time and the active production faces will lag each other by 25 m. The underground mine will commence

with the top three sublevels intersecting the open pit at the 70, 95 and 120 mbsl elevations and the

lowest planned sublevel is at the 970 mbsl elevation.

Underground Mineral Reserve and Schedule

The underground Mineral Reserve Estimate developed for the Rory’s Knoll deposit contains a total of

25.8 Mt of ore at a grade of 2.84 g/t. Underground mining is planned between late 2017 and 2031.

Dilution and recovery modifying factors applied to the estimate are:

• Average LoM dilution grade of 1.13 g/t;

• Average LoM dilution is estimated to be 12%; and

• Average LoM ore recovery is 89%.

The initial construction period will be approximately two years in order to reach commercial production

by 2018. Pre-production development work will be completed by a contractor assisted by owner

personnel and equipment. LoM production will be completed by the owner utilizing its own workforce

and equipment. Upon start-up, production will be at a nominal production rate of 1.9 Mtpa at steady




C O M B I N E D M I N E R A L R E S E R V E E S T I M A T E A N D P R O D U C T I O N1.7S C H E D U L E

The Mineral Reserve Estimate for the Aurora Gold Project has been subdivided into an open pit portion

and an underground portion. Table 1-2 presents the combined open pit and underground reserve

estimate.

Table 1

–

2: Aurora Mineral Reserve Estimate

Quantity Grade Contained Au

Proven (kt) (g/t) (k oz)

OP SAP 168 2.64 14

OP FRESH 2,207 3.07 218

Total Proven 2,375 3.04 232

Probable

OP SAP 4,955 1.70 270

OP FRESH 6,343 3.03 618

Underground 25,851 2.84 2,357

Total Probable 37,149 2.72 3,245

Total P&P 39,524 2.74 3,477

*Open Pit saprolite cut-off grade of 0.3 g/t*Open Pit fresh rock cut-off grade of 0.5 g/t*Underground cut-off grade of 1.2 g/t

The mineral reserve estimate is constrained to a gold price of US$1,300 per ounce, an average




M E T A L L U R G Y 1.8

The objective of the metallurgical feasibility testwork program was to determine the metallurgical

response of the Aurora Gold Project mineralization. The program was designed to develop theparameters for process design criteria for grinding, intensive cyanidation, leaching, cyanide destruction

and determine the solid-liquid response of the ore.

A total of 162 drill core samples representing three zones of the Aurora Gold Project deposit, Rory’s

Knoll, Mad Kiss and Aleck Hill, were submitted to SGS Minerals Services of Lakefield, Ontario. Nineteen

individual zone ore type samples and three composites were prepared for the testwork program. Two

additional composite samples were prepared from samples that were stored at the SGS Lakefield site.

SGS also prepared samples for McGill University Comminution Dynamic Lab. Additional testwork wasperformed by RDi, Inc. of Golden, Colorado on samples of saprolite and fresh rock ore in support of this

updated feasibility study.

The ores tested were highly amenable to cyanide leaching at P80 109 µm with gold recoveries in the 90

to 95 percent range. Grinding finer than P80 109 µm marginally enhanced recovery. Cyanide

consumption from the laboratory tests averaged 0.5 kg/t for fresh rock. This consumption is typical for

a free milling ore with few deleterious cyanide consumers.

Leaching and carbon adsorption kinetic tests indicated there should be no effect of increased pulp

density (in the range of 45% to 55% solids) on either parameter.

The ore is amenable to detoxification by the air/SO2 process (air/SO2/Cu2+), with industry-normal

reagent demand and acceptable weak acid dissociable cyanide (CNWAD) levels for discharge into a wet

tailings dam.




Design availability of 350 days per year (after ramp-up), which equates to 8,400 operating

hours per year, with standby equipment in critical areas;

Sufficient plant design flexibility for treatment of all ore types at design throughput;

Overall gold recovery in excess of 97% in saprolitic ore and in excess of 94% in fresh rock.

PR O J E C T I N F R A S T R U C T U R E 1.10

The remote location of the project site and the absence of local infrastructure dictates that the Project

design incorporate all of the infrastructure components that a large scale mining project requires.

Access to the Project during the exploration phase has been by a 150 km gravel access road from the

Buckhall Port facility, the Cuyuni River and a limited use airstrip. The Buckhall Port facility will be

upgraded to permit ocean going vessels to be docked and will provide facilities for cargo, fuel and

personnel handling during project construction and operations.

At the project site the following principal operations support infrastructure facilities will be constructed:

On-site service roads and heavy equipment haulage roads; Camp accommodation with associated facilities to provide potable water and sewage

disposal will be provided for the accommodation and feeding of construction and

operations personnel;

A No. 4 fuel oil power station and fuel storage facilities for operational power. Power will be

distributed throughout the site by means of overhead lines;




S O C I A L A N D E N V I R O N M E N T A L A S P E C T S 1.11

Environmental Studies

Since 2006, the Aurora Gold Project area of influence has been the subject of several environmental

baseline studies by qualified national and international experts. Initial environmental and social impact

assessments (ESIAs) to Guyana Environmental Protection Agency (EPA) and World Bank

Group/International Finance Corporation (IFC) standards were also conducted in 2009 and 2010. Both

ESIAs are currently being updated to address the updated project design requirements reflected in this

National Instrument 43-101 technical report.

The project is located in north-western Guyana. The climate is tropical, with two distinct wet and twodistinct dry seasons; rainfall is significant, averaging 2124 mm per year. Despite its remote tropical

location and very low population density, the region has been significantly impacted by artisanal and

small-scale mining (ASM) for well over 100 years, as well as by hunting, large-scale logging, and other

intrusive human activities.

The specific location of the project was first explored in the 1930s, and has been impacted by ASM

activities ever since. Environmental studies have therefore largely concentrated on biodiversity and

water quality issues. Species that are indicative of minimally impacted environmental conditions in otherareas of Guyana, such as howler and capuchin monkeys, giant river otters, and harpy eagles, have been

observed to be largely withdrawn or absent from the project’s area of influence, likely due to long-term

ASM, logging, noise, and other effects of human intrusion.

Apart from a single observation of two harpy eagles and one giant otter in 2009, no endangered species

of plants or animals have been observed in the project area over the last six years; the giant otter

observation prompted a special environmental study by a noted specialist that concluded that the




The ESMS is developed to meet current IFC environmental and social performance standards, the

International Cyanide Management Code, and other international best management practices, and is

designed with robust change management processes that will allow it to operate effectively over the life

of the mine. The ESMS and its supporting management plans will also establish requirements and

procedures for a comprehensive environmental monitoring program that considers:

Stability and pH of waste rock/overburden stockpile runoff, to detect the development of

potential acid rock drainage conditions;

Integrity and geotechnical stability of the TMA, as well as seepage and reclaim water quality;

Detection and mitigation of erosion issues that may occur in disturbed areas and

constructed earthworks;

Maintenance of a water balance for the TMA, fresh water pond (FWP) and mine water pond

(MWP), as well as WMP seepage and discharge water quality;

All aspects of the procurement, transportation, and operational/closure phase management

of sodium cyanide reagent;

Delivery, storage, and management of fuel and other hazardous materials;

Spill prevention, control, and contingency planning;

Ambient air quality, noise, and vibration;

Periodic evaluations of biodiversity in the area immediately affected by the project;

Management of hazardous and nonhazardous wastes;




Permit to operation municipal solid waste landfills at the Aurora and Buckhall sites

(Guyana EPA, Ministry of Health, and Central Housing and Planning Authority).

Social or Community Impact

There are no formal or established communities or settlements in the immediate vicinity of the Aurora

site, and the project is not expected to generate the direct socio-economic effects that characterize

many mining projects. It should be noted that:

There are currently no permanent communities or residences within the project

concession that would require any physical displacement or resettlement actions;

There are no known archaeological sites or areas of significant cultural interest

within the project concession; however, as the Guyana National Trust and Ministry

of Culture have expressed an interest in any artifacts or items of potential

historical, archaeological, or anthropological interest that may be encountered over

the life of the project, the Guyana Goldfields ESMS therefore invokes specific

procedures for documenting, protecting, and reporting chance finds;

Implementation of the Community Relations Management Plan for the project will

also provide the means of detecting and appropriately responding to any changing

stakeholder views with respect to cultural heritage concerns, as well as

employment or contracting opportunities, health and safety, and other social

considerations;

The project is not located on lands traditionally owned or customarily used by

indigenous peoples (i.e., Amerindians). The nearest Amerindian community is 50

km upstream of the project, and this community is unlikely to be significantly




C A P I T A L A N D C O S T E S T I M A T E S 1.12

Capital Cost Estimates.

Life-of-Mine (LoM) Project capital is summarized in Table 1-3. Initial capital costs are estimated at $205

million. Expansion capital includes; $93 million for underground mine development (ramp, ventilation

raises, etc.) as well as mining equipment, and $27 million for expansion of the process and power plant.

Table 1–3: LoM Project Capital ( 000s)

Capital CostsInitial

(2013 –2014)Expansion

(2015 –2017)Sustaining

(2018 –2031)Total

(LoM)

Open Pit $12,541 $10,995 $10,392$33,928

Underground $0 $92,612 $315,669 $408,281

Process/ Infra. $175,840 $46,123 $22,620 $244,583

Owner's & Closure $16,545 $1,811 $9,000 $27,356

Total $204,926 $151,541 $357,681 $714,148

Addition differences due to rounding

Initial Capital is scheduled for expenditure during the pre-production period. Pre-production represents

a 24-month period commencing Q1 2013 through Q4 2014 and includes an expansion of on- and off-site

activities currently underway. Expansion Capital includes underground mining and process activities to

increase capacity to 10,000 t/d. Sustaining Capital includes expenditures for Underground mine

development in Q4 2015 to end of life (EoL) in year 2031 and for the Open Pit.

Operating Cost Estimates.

LoM operating costs are summarized in Table 1-4. LoM operating costs are estimated at $1.4 billion, or

$34.95/t-milled. Open pit mining will average $2.42/t-moved ($13.68/t-ore). Underground mining will




Table 1–4: LoM Operating Costs

Cost ItemLoM Cost

($000s)Unit Cost$/t-moved

Unit Cost$/t-ore

Unit Cost$/t-milled

Open Pit Mining $186,999 $2.42 $13.68 -

Underground Mining $498,435 - $19.28 -

Processing $544,551 - - $13.78

G&A $151,225 - - $3.83

Operating Costs $1,381,209 $34.95


F I N A N C I A L A N A L Y S I S 1.13Economic results are summarized in Table 1-4. The analysis indicates the following conclusions assuming

no gearing at a gold price of $1,300/oz:

Mine Life: 17 years

Pre-Tax NPV5%: $1,119 million, IRR: 44%

Post-Tax NPV5%: $800 million, IRR: 38%

Payback (Post-Tax): 40 months

Corporate Income Taxes Paid: $509 million

Cash costs (including Royalty): $527/oz-Au

Peak funding of the initial project capital of $205 million: $163 million in year 2014

Table 1–5: Technical-Economic Results

DescriptionLoM Cost

($000s)Unit Cost$/t-milled

Unit Cost$/oz-Au




DescriptionLoM Cost


Unit Cost$/oz-Au

Post-Tax Cash Flow $1,319,247 - -

NPV5% $799,720 - -

IRR 38% - -

Payback (months) 40 - -


C O N C L U S I O N A N D R E C O M M E N D A T I O N S 1.14

Geology and Resources

Exploration work is professionally managed and field procedures generally meet accepted

industry best practices. SRK is of the opinion that the exploration data are sufficiently

reliable to interpret with confidence the boundaries of the gold mineralization and support

evaluation and classification of mineral resources in accordance with generally accepted

CIM “Estimation of Mineral Resource and Mineral Reserve Best Practices” and CIM

“Definition Standards for Mineral Resources and Mineral Reserves” guidelines;

The bulk of the mineral resources are located in Rory’s Knoll, which represents 69% of the

total reported Measured and Indicated mineral resources and 73% of the reported Inferredmineral resources;

The Aurora gold deposit contains a significant mineral resource estimated at 6.54 million

ounces of gold in the Measured and Indicated categories with an additional 1.82 million

ounces of gold in the Inferred category. SRK notes that the mineral resources occupy a small

footprint on the prospecting license.

SRK considers that the mineral resource model documented herein is sufficiently reliable to support




Geotechnical Recommendations

It is recommended that further detailed engineering incorporate the following:

Additional geotechnical drilling and hydrogeological testing to be conducted around the

Rory’s Knoll deposit to help better characterize the heavily foliated sericite schists and the

interbedded volcanic geomechanical domains (for infrastructure placement).

Additional laboratory testing should be conducted on new core, targeting the sericite schist.

A review and optimization of the geotechnical design criteria for the open pits, which

incorporate results of additional drilling and updated geological and geomechanical

domains.

A review of the slope stability analyses using numerical analyses capable of including thefoliation (i.e. UDEC) for risk of toppling in the open pit.

A recalibration of the geotechnical underground numerical model to incorporate possible

changes to the rockmass characterization and groundwater flow to evaluate the potential

impact on the selected base case.

A detailed geotechnical investigation should be undertaken for the design of the portal or

box cut for underground access.

Development of a detailed open pit and underground instrumentation and monitoringprogram.

Conduction of in-situ stress testing during the pre-development and production

underground phases.

Conduction of studies to evaluate the potential for the risk of mudrush and develop

standard operating procedure for such events.

Estimated budget for the open pit evaluation is US$50,000. Estimated budget for the underground




Open Pit Mining

The near surface mineralization at the Aurora Gold Project is amenable to conventional

loader/truck mining methods utilizing 7.7m3 front end loaders and 43.5 tonne class

articulated trucks.

Financial modeling of the open pit has determined that the open pit is economically viable

and supports Proven and Probable Reserves. The open pit Reserves are 13.7 Mt of ore

grading 2.55g/t gold. This reserve includes 2.4Mt of ore in the Proven category.

Open pit development includes haul road construction and pre-stripping in 2014 with millproduction beginning in 2015.

The open pit will feed the mill 1.75M tonnes per year.

The total mining rate for the open pits will average 20,700t/d over the life of mine. The

average is 14,200t/d over the first three years of mining. The average is 31,400t/d over the

last six years of the mine life.

The open pit will require US$12.5M of initial pre-production capital. Total capital cost of the

life of mine will be US$33.9M.

The average operating cost of the open pit will be US$2.42 per tonne mined. The initial unitcost will be US$3.19 over the first three years of operation.

It is recommended that further detailed engineering incorporate the following:

Dewatering sump and ditch locations.

Evaluation of pit designs for zones excluded from the feasibility study.

Evaluation of diesel versus electric pumps for pit dewatering.




Underground mining includes portal construction and capital decline development

commencing in the fourth quarter of 2015, with underground commercial production

beginning in early 2018.

The underground mine will feed the mill at a nominal rate of 1.9M tonnes per year.

The underground mine will require US$92.6M of initial pre-production capital and

US$315.7M of sustaining capital.

The average operating cost of the underground mine is US$19.28 per ore tonne mined.

A comprehensive underground geotechnical instrumentation and monitoring program has

to be implemented to mitigate potential risk of larger than expected stope wall failures

(refer to the technical risks section).

A comprehensive training program will be required to train local labour for the undergroundmine.

It is recommended that further detailed engineering studies incorporate the following:

Performance of a tradeoff study to evaluate the Rail-Veyor material haulage system as an

alternative to conventional truck haulage.

Performance of a scoping study to evaluate an amendable and economical underground

mining method to extract additional Mineral Resources below the satellite pits. Performance of a scoping study to evaluate the optimum underground mining depth by the

SLR mining methods.

Estimated budget for this work is US$300,000.

Mineral Processing

The testwork indicated that the Aurora ores are amenable to conventional processing




Detailed engineering will be required to prepare construction-ready documents and to

finalize construction cost estimates. Estimated budget for this work is $3,000,000;

Based on the investigation, analysis and design work carried out by AMEC, some areas of theproject still require further investigation to be carried out, regarding hydrogeology (pump

test and packer test along the river dike) and geotechnical work (at some of the dam

locations). The proposed works should be carried out prior to or in parallel with the early

stage of the detailed design. Estimated budget for this work is $100,000;

A trade-off study comparing the use of No. 2 Diesel Fuel versus No. 4 Fuel Oil should be

conducted to determine the most efficient method to produce power. An alternative for

producing power at Buckhall Port with a transmission line to Aurora mine should also be

considered. An additional alternative study of a biomass power plant at either the Auroramine or at Buckhall Port should also be considered. Estimated budget for this work is

$20,000.

Environmental and Social

The project’s area of influence (AOI) has been significantly impacted by historical

artisanal and small-scale mining (ASM), logging, and hunting, for well over a hundred

years;

Large fauna that are otherwise common in pristine habitats along similar types of rivers

in this area of South America are absent or rare in the project AOI, and may be viewed

as a key indicator of significant historical human impact;

With very few exceptions, rare, threatened, or endangered species have not been

observed in the area of the project;

There are no formal or established communities or settlements in the immediate




2 . 0 I N T R O D U C T I O N

B A C K G R O U N D I N F O R M A T I O N 2.1

2.1.1 General Project Description

The Aurora Gold Project is a development stage gold exploration project situated in the forested region

of Guyana, in an uninhabited area, approximately 170 km west of the capital Georgetown. Guyana

Goldfields Inc. (Guyana Goldfields) holds a 100% interest in the project.

Infrastructure and local resources are virtually non-existent in the project area, except for that

constructed by Guyana Goldfields. The property extends southeast from the Cuyuni River, which is

approximately 30 m above sea level. The area is of low relief and covered with dense rainforest. The low

relief results in large swampy areas during the rainy seasons that cover parts of the license area.

Guyana Goldfields has constructed a runway on the southern bank of the Cuyuni River adjacent to the

existing exploration camp, which is suitable for helicopters and short take-off-and-landing aircraft.

Access by air using this runway is currently the safest and most expedient method for personnel andsmall equipment parts to access the property. However, the primary site access is by road from Buckhall

Port on a maintained gravel road and barge crossing at Tapir Crossing. Road egress is 150 km from

Buckhall Port to the Aurora Project site with overland travel taking approximately six hours.

2.1.2 Terms of Reference

In May 2012, Tetra Tech, Inc. (Tetra Tech) was retained by Guyana Goldfields to prepare this Updated

Feasibility Study. The purpose of updating the previous Feasibility Study was to investigate opportunities




Under a separate commission with Guyana Goldfields, SRK prepared the resource block model, open pit

mine planning and related cost estimation; underground mine planning and related cost estimation;

used in this Feasibility Study. Results of that resource estimate were publicly released by Guyana

Goldfields on June 25, 2012.




3 . 0 R E L I A N C E O N O T H E R E X P E R T S

Tetra Tech relied on RDi, Inc. for metallurgical review and process considerations.

SRK relied on BBE, Inc. for evaluation of requirements for underground ventilation and cooling (refer to

Sections 16.3.16, 16.3.17). SRK relied on Itasca, Inc. for evaluation and modeling of groundwater

hydrogeology (refer to Section 16.1).

Tetra Tech relied on the expertise of Mr. Paul Murphy, Executive VP, Finance and CFO, Guyana

Goldfields in setting up the correct functioning of the taxes section of the feasibility study financial

model. Tax matters are addressed in the “Aurora Mineral Agreement,” of November 18, 2011, between

the Co-operative Republic of Guyana and Guyana Goldfields.

A review of land title and tenure was prepared by Jonas M.F. Coddett & Associates, Attorney-at-Law,

legal advisor to Guyana Goldfields on March 4, 2012.

Mr. Paul Murphy (Executive VP, Finance and CFO) indicated that there are no known litigations that

would potentially affect the Aurora Gold Project.

Tetra Tech relied on the Feasibility Study Global & Local Logistics & Cargo Shipping Services from Asia,

Europe & North America for Guyana Goldfields, Inc. prepared by Fracht Canada Freight, Inc. dated

October 9, 2012 for evaluation of project shipping and logistics.




4 . 0 P R O P E RT Y D E S C R I P T I O N A N D

L O C A T I O N

The Aurora Gold Project is located in Guyana, South America, approximately 170 km west of the capital

Georgetown and 130 km west north-west of Bartica, a settlement at the junction of the Essequibo and

Cuyuni Rivers. Bartica is a regional hub for accessing the interior of north-western Guyana. The center of

the property is located at latitude 6°45′N, longitude 59°45′W (Figure 4-1). The project includes the

Buckhall Port on the Essequibo River. There is a 150 km road from Buckhall Port to the Aurora Project

site, and a ferry crossing of the Cuyuni River at Tapir.

The general area of the Aurora Gold Project has been subject to mineral exploration since the 1940s.

This part of Guyana is largely uninhabited with the nearest settlement approximately 50 km away.




The boundaries of the A1 license prospecting license form trends approximately southeast-northwest,

south of the Cuyuni River. The northern edge of the shape follows the south bank of the Cuyuni River; all

other edges are straight and are defined by six corner points, which are listed in Table 4 –1.

Table 4–1: Corner Points of Prospecting License A1

Corner Point ID Latitude Longitude

A 6°47'32"N -59°43'17"W

B 6°45'38"N -59°41'24"W

C 6°43'03"N -59°41'24"W

D 6°43'03"N -59°43'08"W

E 6°46'40"N -59°46'54"W

F 6°48'15"N -59°46'54"W

Guyana Goldfields has confirmed that the mineral tenure, surface rights as well as access and permitting

issues of the Aurora Gold Project have been reviewed and were found to be in good standing by

independent legal counsel (Appendix A-letter from Guyana Goldfields attorney).

The Aurora Gold Project is located in a remote part of the rainforest; hence, a precise description of the

property boundaries is difficult. Guyana Goldfields retained Edward Luckhoo (from Montejo, 2009), a

registered lawyer, to supply a legal opinion on the land position. Mr. Luckhoo’s opinion is:

“From a reference point ‘X’ located with geographical co-ordinates of latitude 6°47′10″N, longitude

59°42′05″W and situated at the confluence of the Cuyuni River and Gold River, thence going upriver for

a distance of approximately 1 mile 704 yards to point of commence ‘A,’ located with true geographical

co-ordinates of latitude 6°47′32″N, longitude 59°43′17″W; thence at a true bearing of 135° for a distance







originally required to make annual advance royalty payments to Mr. Alphonso in the aggregate of

US$225,000 per year during the three year period following the commencement of commercial

production, and to pay an additional 2% net smelter royalty (NSR) to Mr. Alphonso thereafter. On March

18, 2004, the original agreement was amended, pursuant to which Guyana Goldfields agreed to pay Mr.

Alphonso an annual fee of US$100,000 for as long as Guyana Goldfields maintains an interest in the

Aurora Gold Project, up to a maximum of US$1,500,000.

PE R M I T S A N D A U T H O R I Z A T I O N S 4.3

All exploration programs to date were conducted under appropriate authorization, license, or equivalent

control documents, which were obtained from the appropriate regulatory authority in Guyana.

The Mining Licence, obtained in November 2011, gives the company the right to build and operate the

mine. At the same time the company signed a Mineral Agreement (MA) with the Government of Guyana

and the Guyana Geology and Mines Commission which sets the fiscal regime, taxation and royalties as

they effect the operation of the mine. This licence and the MA were signed by the Company and the

Company’s wholly owned subsidiary in Guyana, AGM, Inc. and are valid for 20 years and renewable on

application for further 7 years periods for as long as mining operations continue on the property.

Significant details among the MA terms include:

Mining royalty of 5% on gold sales at a price of gold of US$1,000/oz or less;

Mining royalty of 8% on gold sales at a price of gold over US$1,000/oz;

Corporate income tax rate of 30% and no withholding tax on interest payments to lenders;

Duty and value added tax exemptions on all imports of equipment and materials for all

continuing operations at the Aurora Gold Project, including the construction and operation




5 . 0 A C C E S S I B I L I T Y, C L I M A T E , LO C A L

R E S O U R C E S , I N F R A S T R U C T U R E A N DP H Y S I O G R A P H Y

A C C E S S I B I L I T Y 5.1

5.1.1 Accessibility by Air

Guyana has two international airports. Cheddi Jagan International Airport is approximately one hoursouth of Georgetown, the nation’s capital and the airport is serviced by international carriers. A smaller

national and limited international airport, Ogle Airport, located 10 km east of Georgetown, provides

access to regional Guyana and adjacent countries.

Guyana Goldfields has constructed a runway on the southern bank of the Cuyuni River adjacent to the

camp, which is suitable for helicopters and short-takeoff-and-landing aircraft. The runway has an

approximate length of 700 m. Guyana Goldfields operates several charter flights per week from Ogle

airport to the project site.

5.1.2 Ground Accessibility

The Aurora Gold Project is also accessed via a road from the Buckhall Port facility. The road distance

between the Buckhall Port facility and the Aurora Gold Project site is approximately 150 km. The road

alignment initially follows the north shore on the Cuyuni River, and crosses over the river at the Tapir

Crossing via barge to continue to the mine site. Much of the existing road was constructed and is

currently being maintained by Barama Company Limited for logging. Barama will continue to use their




There are two wet seasons, from April to August and from December to January, and two dry seasons,

from February to April and from August to December. Average rainfall in the forest region is 2,124 mm

per annum.

Relative humidity is high, ranging from 65% to 100%. Temperatures range from 22° C to 34 °C year

round.

The humid tropical climate of Guyana is moderated by the north-eastern trade winds. Exploration

activities and mining operations will be conducted year-round.

LO C A L R E S O U R C E S A N D I N F R A S T R U C T U R E 5.3

The project is located in a very remote and uninhabited area of Guyana. Project execution will require

building all required infrastructure a portion of which has already been constructed. The access road to

the project has been completed. The Aurora Gold Project has an existing mancamp, light maintenance

and fuel storage facilities, and an expanded camp to accommodate construction activities is underway.

Basic supplies are available in Georgetown, which has a population of approximately 240,000. The city is

located approximately 40 km east of Buckhall Port.

Most major items and equipment will be imported from overseas. Access to the site for project

development will be primarily by road or air. Equipment and supplies entering the site will clear customs

at Buckhall Port.

Power for future mining operations will be generated on-site. The power plan calls for the use of on-site

generator sets using No. 4 Fuel Oil. Guyana Goldfields is currently undertaking a study to determine the

feasibility of alternative energy sources for generating electric power.







Small hills are also present to the southwest of the property and rise approximately 40 m above river

level. These hills are formed of granitic rocks and clay-rich residual deposits that are cut by streams that

drain into the Cuyuni River.




6 . 0 H I S T O R Y

E X P L O R A T I O N W O R K P R I O R T O G U Y A N A G O L D F I E L D S 6.1

Details about legacy exploration programs are limited. Information given in this section is sourced from

Cargill and Gow (2003) and Cargill (2005), and is summarized in Table 6 –1.

Table 6–1: Summary of Historical Exploration Work Prior to 1998

Period Company Activity Drilling UG Development Production

1911 Discovery of gold.

1934 -

1937

Numerous claims staked.

1938 -

1939

Solar Exploration.

1940 -

1948

Cuyuni Systematic development of

claims, mining started in 1940.

30 surface

(4,809 m)

26 UG

(1,600 m)

To depth of ~ 75

m below surface

at Aleck Hill

Est. 2,260 -

3,800 kg Au

1963 Geol. Survey

of Guyana

Geochemical and geophysical

surveys.

19 surface

(2,515 m)




In 1940, Cuyuni Goldfields Company (Cuyuni) acquired the rights to part of the project area and began

to develop their mineral claims systematically. In 1945, Cuyuni was able to acquire the remainder of the

claims that comprise the current project area. Mining activities commenced in 1940 and continued until

1948, at which point underground development at Aleck Hill had reached a depth of approximately 75

m below surface. Mining records are either missing or lack detail; hence, any production figures that are

available are estimates at best. Webber (1952) estimated that approximately 2,260 to 3,800 kg of gold

were produced by Cuyuni from mineralization with an average head grade of approximately 18 g/t Au.

Cuyuni drilled 26 surface core boreholes (4,321 m) and 26 underground core boreholes (1,600 m) at

Aleck Hill and 4 surface core boreholes (488 m) at Mad Kiss.

Cuyuni ceased mining operations in 1948 and the project area lay dormant until 1963, when the

Geological Survey of Guyana conducted an exploration program in the Haimaralli Falls area, along the

northwest border of the Aurora project area. This program was aimed at identifying copper

mineralization. The Geological Survey of Guyana carried out geochemical and geophysical surveys,

consisting of Turam electromagnetics and ground magnetics, and completed 19 core boreholes (2,515

m). No significant copper mineralization was intersected.

No exploration work was carried out in or around the project area between 1963 and 1989. In 1989,

South American Goldfields Inc. (South American) acquired an Exclusive Exploration Permit covering the

Aurora project area. South American did not carry out any exploration work but had an agreement with

Denison Mines Ltd. (Denison) to carry out exploration. Commencing in 1989, Denison completed a three

year exploration program comprising gridding, soil, rock chip, saprolite, and stream sediment sampling;

geological surface and underground mapping; underground sampling; and acquisition of airborne and

ground geophysical data. Denison also drilled 56 core boreholes (10,204 m).

The aeromagnetic survey, carried out by Denison in 1990, covered the entire Aurora Project area. Initial




Sometime during the mid-1990s, Mr. Alfro Alphonso acquired the property and subsequently optioned

the property to Coeur d’Alene Mines Ltd., who carried out a geochemical exploration program.

E X P L O R A T I O N B Y G U Y A N A G O L D F I E L D S (1998 T O 2009)6.2

During 1998, Guyana Goldfields acquired a 100% option on the property from Mr. Alfro Alphonso. An

unknown amount of geological mapping has been completed on the project area.

In 2002 and 2003, Guyana Goldfields conducted a drill program comprising 39 shallow core boreholes

(1,076 m), deep auger sampling, trenching, and channel sampling on the A1 License. In December 2004,

Guyana Goldfields obtained a reconnaissance permit covering approximately 600 km 2 surrounding the

original A1 License.

Airborne magnetic, radiometric and electromagnetic surveys, and trenching and channel sampling were

completed in 2004. Based on this exploration data, Guyana Goldfields applied for five new prospecting

licenses contiguous with the original A1 License. These licenses were formally granted by the

Government of Guyana on June 29, 2004.

From 2004 to 2009, delineation drilling was completed at the Aleck Hill, Rory’s Knoll, Walcott Hill,

Aurora and Mad Kiss areas. A total of 851 boreholes were drilled (196,301 m). A petrography study wasalso completed in 2005 (Kipfel, 2005) and an independent structural study by SRK was completed in

April 2007.

A Mineral Resource Statement was prepared by Micon for Guyana Goldfields in November 2007

(Mukhopadhyay, 2007). During 2008, a Preliminary Economic Assessment (PEA) was prepared by

Snowden Associates (Myer, 2008).




Table 6–2: Summary of Exploration Work by Guyana Goldfields between 1998 and 2009

Period Company Activity Drilling

1998-2009 GuyanaGoldfields

Geological mapping, geophysical surveys, geochemicalsampling, trenching, drilling, petrography, independentstructural study, mineral resource estimates, preliminaryassessments, technical reports.

890 holes(197,377m)

PR E V I O U S M I N E R A L R E S O U R C E E S T I M A T E S 6.3

Micon prepared the first two Mineral Resource Statements for the Aurora Project published in

November 2007 and December 2008. The third Mineral Resource Statement was prepared by AMEC in

June 2009, and was considered for a preliminary economic assessment also by AMEC.

The AMEC (Montejo et al, 2009) Mineral Resource Statement was based on exploration data to March

30, 2009 (508 core boreholes). The statement considered 173 resource domains generated by Guyana

Goldfields.

Open pit mineral resources were constrained by a conceptual pit and underground mineral resources

were reported below the conceptual pit. The conceptual pit envelope was designed at a gold price ofUS$750/oz. The mineral resources were reported at a range of cut-off grades, with the base case

reported at cut-off grades of 0.85 g/t Au for open pit and 2.00 g/t Au for underground mineral

resources, respectively (Table 6 –3).

Table 6–3: Mineral Resource Statement, Aurora Gold Project, AMEC, June 2, 2009

Classification Quantity Grade Contained Au

000’ Tonnes Au (g/t) 000’ Ounces




Table 6–4: Mineral Resource Statement, Aurora Gold Project, SRK Consulting (Canada) Inc., February 28,

2011

Classification Quantity Grade Contained Au

000’ Tonnes Au (g/t) 000’ OuncesOpen Pit

Measured 5.59 3.44 0.62Indicated 11.69 3.54 1.33Measured and Indicated 17.28 3.51 1.95Inferred 3.53 3.74 0.42

UndergroundMeasuredIndicated 24.89 4.25 3.40Measured and Indicated 24.89 4.25 3.40Inferred 6.90 4.10 0.91

Combined MiningMeasured 5.59 3.44 0.62Indicated 36.58 4.02 4.73Measured and Indicated 42.17 3.94 5.35Inferred 10.43 3.98 1.33

* Mineral resources are not mineral reserves and do not have demonstrated economicviability. All figures have been rounded to reflect the relative accuracy of the estimates. The

cut-off grades are based on a gold price of US$1,045/oz and metallurgical recoveries of95% for saprolite and fresh material. Open pit mineral resources are reported at a cut-offgrade of 0.45 g/t Au inside conceptual pit shells, whereas underground mineral resourcesare reported at a cut-off of 2.0 g/t Au.

SRK prepared the fifth Mineral Resource Statements, which was published in September 2011. The

borehole database contains updated drilling data for the period December 2010 to May 2011. It

considered 939 exploration boreholes (291,556 m; excluding geotechnical and metallurgical holes), and

67,843 gold assay intervals.




Table 6–5: Consolidated Mineral Resource Statement* Aurora Gold Project, Guyana, SRK Consulting

(Canada) Inc., September 9, 2011

ClassificationQuantity Grade Contained Au

000’ Tonnes Au (g/t) 000’ OuncesOpen Pit Mining Measured 5.75 3.29 0.61Indicated 14.47 3.31 1.57Inferred 3.48 3.41 0.39Underground MiningMeasured 0 0 0Indicated 26.82 4.09 3.52Inferred 6.49 3.74 0.78

Combined MiningMeasured 5.75 3.29 0.61Indicated 41.29 3.83 5.10Inferred 9.97 3.63 1.17

* Mineral resources are not mineral reserves and do not have demonstrated economicviability. All figures have been rounded to reflect the relative accuracy of the estimates. Thecut-off grades are based on a gold price of US$1,200 per ounce of gold and metallurgicalrecoveries of 95% for saprolite and fresh material. Open pit resources are reported at a cut-off grade of 0.40 g/t Au within conceptual pit shells. Underground mineral resources arereported at a cut-off grade of 1.8 g/t Au and include all blocks above cut-off outside the

conceptual pit shells.




7 . 0 G E O L O G I C AL S E T T I N G A N D

M I N E R A L I Z A T I O N

R E G I O N A L G E O L O G Y 7.1

The Aurora Gold Project is located in the Archean-Proterozoic Guiana Shield in northeast South America.

The Guiana Shield is a palaeo-Proterozoic granite-greenstone terrane and is considered to be the

extension of the West-African palaeo-Proterozoic Birimian Supergroup terrane. The Guiana Shield islargely composed of the Barama-Mazaruni Supergroup, a metasedimentary/greenstone terrane

intercalated with Archean-Proterozoic gneisses that are intruded by Trans-Amazonian granites, as well

as mafic and ultramafic rocks (McConnell and Williams, 1969).

The Barama Group consists of pelitic metasedimentary and metavolcanic rocks. The Mazaruni Group

conformably overlies the Barama Group, which also consists of metasedimentary and metavolcanic

rocks. The Mazaruni Group is subdivided into the Cuyuni Formation and the Haimaraka Formation.

The Cuyuni Formation consists of pebbly sandstone and intraformational conglomerate, intercalated

with felsic to mafic volcanic rock. The Haimaraka Formation conformably overlies the Cuyuni Formation

and consists of a thick sequence of mudstone, pelite, and graywacke; significant amounts of volcanic

rock are absent from this unit (McConnell and Williams, 1969).

The Barama-Mazaruni Supergoup formed within a geosynclinal basin locally bordered by an Archean

continental foreland. The Trans-Amazonian Orogeny, approximately 2 Ga, resulted in block faulting,

crustal shortening folding metamorphism and anatexis of the Barama-Mazaruni Supergroup (Hurley et










The locally named tonalite rock at Rory’s Knoll is in actuality a mafic rock, possibly dioritic in composition

that has undergone intense hydrothermal alteration including silicification. It is unclear if this distinctive

rock unit is intrusive. It is a competent lithology in which auriferous quartz-ankerite veins represent

dilational sites developed during active deformation and fracturing. The auriferous veins of Walcott Hill,

Walcott Hill East and Rory’s Knoll East are hosted within a similar rock type.

The quartz and feldspar porphyry (QFP) dike(s) modeled in the Mad Kiss area, which could be referred to

as a quartz phyric felsic intrusion, is also a competent lithology in which auriferous veins formed in

response to dilation of that stiff rock unit.

Metasedimentary rocks logged as either “Ash Tuff” or “Metasediment” in the drill database consist of

turbiditic laminated to thickly-bedded argillite and greywacke.

The interbedded sequence of metasedimentary and mafic metavolcanic rocks consists primarily of

metasedimentary rock with mafic metavolcanic rock subunits. The interbedded unit is especially

prevalent northeast of the Rory’s Knoll deposit.

Mafic metavolcanic rocks form the southwest contact of Rory’s Knoll, the host rock of the Aleck Hill

deposit, and the Mad Kiss Quartz and Feldspar Porphyry. Some mafic rocks are very strongly magnetic,

especially in the Mad Kiss and Walcott Hill areas. This is well emphasized on the aeromagnetic data.

M I N E R A L I Z A T I O N 7.3

Gold mineralization in the Golden Square Mile area is controlled by a series of northwest-southeast

trending shear zones. These shear zones are orientated sub parallel to the dominant northwest-

southeast structural trend that occurs throughout the Aurora property. The shear zones contain a steep

northwest-southeast trending foliation that formed during northeast-southwest shortening.







The steep northerly plunge at Rory’s Knoll is interpreted to represent the intersection between a sub

vertical northwest trending shear zone and a steeply dipping, possibly west trending stiff lithology

(altered diorite). Gold mineralization at Rory’s Knoll persists to the west in Walcott Hill East.

122.0m

131.1m

134.8m

136.0m

138.9m

140.2m

142.2m

145.5m

149.2m

157.0m

159.8m

132.3m

122.0m

131.1m

134.8m

136.0m

138.9m

140.2m

142.2m

132.3m

Barren “Tonalite

Rock”

Sheared contact

with Qz – Cb

veining




better delineated by drilling. The Walcott Hill East Zone has been included in the drilling and wireframes

for Rory’s Knoll.

Source: SRK, 2011 (Figure provided by Guyana Goldfields)

Figure 7–5: Typical Section through Walcott Hill East (A. Borehole EWD-19 & B. Borehole EWD-6)

7.3.2 Aleck Hill

Aleck Hill is located approximately 1,000 m southwest of the Rory’s Knoll Zone . Gold mineralization





Figure 7–6: Typical Section through Aleck Hill (Borehole AHD-7)

At Aleck Hill North, high strain and alteration zones developed at gradational contacts between diorite

and mafic volcanic rocks (Figure 7-7).







7.3.3 Walcott Hill

Walcott Hill is located approximately 500 m southwest of the Rory’s Knoll Zone (Figure 7-3).

Unpublished exploration reports from Guyana Goldfields describe the gold mineralization at Walcott Hillas gold-bearing quartz veins with a thickness of up to 0.6 m striking 320° and dipping sub vertically for

about 35 m along strike. The width is uncertain and the extent along strike and dip is currently poorly

defined. The Walcott Hill Zone has been intersected by drilling to a depth of approximately 650 m below

surface. Forty-seven wireframes have been modeled for Walcott Hill.

7.3.4 Mad Kiss

The Mad Kiss Zone is located approximately 750 m south-southwest of the Rory’s Knoll Zone (Figure 7-

3). In the Mad Kiss Zone, quartz-carbonate veining occurs inside a sheared quartz-feldspar porphyry dikeenclosed in foliated muscovite-rich rock (Figure 7-8). The gold-bearing stockwork system trending about

150° and dipping steeply north and south is also associated with the hanging wall and footwall contacts

of the porphyry dike. The quartz feldspar porphyry dike is up to 150 m wide. Gold mineralization trends

250° dipping 70° north. Auriferous veins are 2 cm to 5 cm in thickness and occur parallel and normal to

the regional foliation. The lower contact of the quartz feldspar porphyry dike is sometimes marked by a

thick quartz-carbonate vein with variable gold grades. There is no veining in muscovite-rich rock.








Figure 7

–

9: Typical Section through Mad Kiss West (Borehole WMKD-16)




8 . 0 D E P O S I T T Y P E S

The gold mineralization at the Aurora Gold Project exhibits features analogous to mesothermal or

“orogenic” gold deposits typified by Archaean deposits of the Abitibi region, Canada. Features

characteristic of the gold mineralization at the Aurora Gold Project include:

Relative late timing during active compressional deformation;

A strong spatial association to large scale shear zones;

Formed during greenschist metamorphic conditions;

Association with a propylitic-phyllic alteration assemblages; and

Is principally hosted in quartz-ankerite-pyrite veining.




9 . 0 E X P L O R A T I O N

The exploration work conducted prior to 2009 is described briefly in Section 6 and more details about

the exploration work completed by Guyana Goldfields in particular can be found in the AMEC 2009

Technical Report (Montejo et al., 2009) and is not repeated here.

This report presents the exploration work carried out by Guyana Goldfields from December 1, 2010 to

April 30, 2012, the cut-off date for these data considered for the Mineral Resource Statement

documented herein. Exploration undertaken since December 2010 included primarily core drilling toexpand and infill the various gold zones identified within the Golden Square Mile area (see Section 6).

Diamond drilling (except geotech) done at Aurora Gold Project after the resource cut-off date:

Table 9–1: Diamond Drilling After the Cut-off Date

Prospect No. of Drill Holes Total Meters

Sand Creek 5 908

Gold Creek 5 813

Condemnation 5 1,345

15 3,066




1 0 . 0 D R I L L I N G

Drilling information prior to 2009 was summarized from Montejo et al. (2009). Information on drilling

programs prior to Guyana Goldfields involvement is limited; available data includes 131 core boreholes

(19,128 m) drilled by Cuyuni, the Geological Survey of Guyana, and Denison. Table 10 –1 shows a

summary of those drilling programs. Historical drilling is also discussed in Section 6.0 of this report. Data

from these historical boreholes were not considered for resource estimation.

Table 10–1: Summary of Historical Drilling on Aurora Gold Project

Company Year AreaNumber ofBoreholes

Length(m)

Cuyuni 1940-1948 Aleck Hill Surface 26 4,321

Aleck Hill U/G 26 1,600

Mad Kiss 4 488Subtotal 56 6,409

Geological Survey ofGuyana

1963 Haimaralli Falls 19 2,515Subtotal 19 2,515

Denison 1989-1991 Aleck Hill 22 4,550

Aleck Hill South 2 405

Mad Kiss 16 2,233

Mad Kiss South 10 1,850

Walcott Hill East 3 552

Walcott Hill 1 286

Aleck Hill North 1 205

Mad Kiss West 1 123

Subtotal 56 10,204

Total 131 19,128

3 101




Table 10–2: Summary of Drilling by Guyana Goldfields Inc. between 2002 and 2012

Year AreaNumber ofBoreholes

Length(m)

2002 – 2003

Aleck Hill 26 738Mad Kiss 9 213Felice 4 125Subtotal 39 1,076

2004 – 2005

Aleck Hill Fresh Rock 15 3,103 Aleck Hill Saprolite 26 1,956 Aleck Hill North 5 1,468Walcott Hill East 6 1,264Walcott Hill 11 1,681Mad Kiss South 8 825

Mad Kiss 14 2,672South East Aurora 5 1,161Rory’s Knoll 52 17,462Subtotal 142 31,592

2006

Aleck Hill Fresh Rock 26 7,187 Aleck Hill Saprolite 18 1,266Rory’s Knoll 28 12,400 Aleck Hill North 9 1,952 Walcott Hill East 11 3,744Mad Kiss 6 1,450Mad Kiss West 18 5,941South East Aurora 3 978 Felice 2 449Geophysical Anomalies 11 2,814Port-knockers Workings 11 1,997Swamp Vein 6 1,017Powis Hill 5 813Marupa 13 3,009Subtotal 167 45,017 Aleck Hill Fresh Rock 11 3,656Aleck Hill Saprolite 11 778

NI 43 101 T h i l R t




Year AreaNumber ofBoreholes

Length(m)

Mad Kiss South 17 2,387Mad Kiss West 5 705Swamp Vein 8 1,345Rock Mechanics 18 5,716

Soil Geotechnical 41 2,116

Metallurgical 7 1,011 Condemnation 13 2,684 Subtotal 292 50,844

2010

Aleck Hill 72 24,010 Rory’s Knoll 14 3,093 Aleck Hill North 14 3,424Walcott Hill East 12 2,402Mad Kiss 58 20,245Mad Kiss West 4 1,717Powis Hill 6 2,153Condemnation 25 7,586Rock Mechanics 10 2,256Soil Geotechnical 8 2,007Subtotal 223 68,893 Aleck Hill 120 40,051Rory’s Knoll 20 12,751 Aleck Hill North 46 11,226

Walcott Hill East 38 10,8782011 Mad Kiss 37 15,546

Mad Kiss West 32 8,299Mad Kiss South 16 3,198Marupa 13 1,613Condemnation 21 6,721Rock Mechanics 3 2,051Soil Geotechnical 28 2,051Subtotal 374 114,385 Aleck Hill 7 4,012

NI 43 101 T h i l R t




NI 43 101 T h i l R t




D R I L L C O R E S A M P L I N G 10.3

From July 2009, five geologists were assigned to the project to ensure orderly monitoring of the drilling

program. One geologist was assigned to quality assurance and quality control and all core sampling wasconducted under his supervision.

After drilling the core was placed in plastic core boxes at the drill rig holding three m of HQ and/or NQ

diameter sized core (6.35 and 4.76 cm diameter, respectively). Core boxes were then transported to the

Aurora Camp for logging and sampling. Drill core is stored on the property in plastic core boxes.

The core was photographed and rock quality designation (RQD) measured. Logging was carried out by

Guyana Goldfields geological personnel recording lithology, alteration, mineralization and structuralfeatures of the core. Once logging was completed, sulphide mineralized, altered and quartz veined

sections were marked for sampling. Both bedrock and saprolite core were sampled. Core recovery is

very good, usually approximating 100%, except locally in strongly muscovite-altered rocks.

Sample length is based on geology and sample intervals do not cross lithological contacts. Sample

lengths range from 1 to 3 m.

Unweathered samples were cut in half using a diamond saw and saprolite core was usually cut in half

with a knife with fragments of quartz vein material split in a Longyear core splitter.

Assay samples were labeled and placed into a plastic sample bag and sealed for shipment to the Acme

Laboratory sample preparation facility in East Coast Demerara, Guyana. The other half of the core

sample was returned to the core box with the sample interval and sample number clearly indicated on

the core box with the split core.

Once a sufficient amount of samples were collected, the samples were delivered to the Acme

NI 43 101 Technical Report




1 1 . 0 S A M P L E P R E P A R AT I O N , A N A LY S E S , A N D

S E C U R I T Y

Sample preparation, analysis, and security prior to 2009 are described in detail in Montejo et al. (2009)

and are only summarized below.

Until the 2009 drill program, Guyana Goldfields sent all but umpire samples to Loring Laboratories

(Guyana) Ltd. (Loring) for sample preparation and assaying. Loring is a small, unaccredited laboratory

with two laboratories, one in Guyana and one in Calgary, Alberta. Between 2004 and 2006, GuyanaGoldfields submitted an unknown number of samples for check assaying to ALS Chemex Laboratories

(ALS) in Santiago, Chile as well as to the Omai gold mine laboratory operated by Cambior Inc. in Guyana

for check assaying. ALS operates under a global quality management system that is accredited

ISO9001:2000. The laboratory facility at the Omai gold mine was not accredited.

In early 2009 AMEC and subsequently Guyana Goldfields conducted extensive reviews of sample

preparation procedures and the analytical performance of the Loring laboratory. Following those very

extensive reviews, Guyana Goldfields hired a quality assurance and control manager and improved theiranalytical procedures.

During the 2009 to 2012 period, Guyana Goldfields has used Acme Analytical Laboratories Ltd. (Acme) in

Georgetown, Guyana and Santiago, Chile as their primary preparation and assaying laboratories. The

management system of both Acme laboratories in Georgetown and Santiago is accredited ISO

9001:2000.





QU A L I T Y A S S U R A N C E A N D Q U A L I T Y C O N T R O L P R O G R A M S 11.2

Quality control measures are typically set in place to ensure the reliability and trustworthiness of

exploration data. These measures include written field procedures and independent verifications ofaspects such as drilling, surveying, sampling and assaying, data management and database integrity.

Appropriate documentation of quality control measures and regular analysis of quality control data are

important as a safeguard for project data and form the basis for the quality assurance program

implemented during exploration.

Analytical control measures typically involve internal and external laboratory control measures

implemented to monitor the precision and accuracy of the sampling, preparation and assaying. They are

also important to prevent sample mix-up and to monitor the voluntary or inadvertent contamination ofsamples.

Assaying protocols typically involve regularly duplicating and replicating assays and inserting quality

control samples to monitor the reliability of assaying results throughout the sampling and assaying

process. Check assaying is normally performed as an additional test of the reliability of assaying results;

it generally involves re-assaying a set number of sample rejects and pulps at a secondary umpire

laboratory.

11.2.1 Quality Assurance and Quality Control Programs Prior to 2009

From 2000 to 2003 Guyana Goldfields did not have formal analytical data quality control procedures,

relying on the quality control measures undertaken by the primary laboratories. Starting in 2004,

Guyana Goldfields began inserting control samples (blank and certified reference material samples)

within samples batches submitted for assaying. Blank material was sourced from crushed coarse granite

near Georgetown, Guyana. Guyana Goldfields purchased 12 different reference material standards from

Ore Research & Exploration Pty Ltd (Ore Research) Bayswater North Victoria Australia with gold





Table 11–1: Specifications of Control Samples used by Guyana Goldfields on the Aurora Gold Project

between November 30, 2010 and July 31, 2011

Standard

ReferenceMaterial

Source

Recommended

Au Value(ppm)

StandardDeviation

Number

ofSamples

CDN-BL-4 CDN Resources <0.01 - 113

CDN-BL-7 CDN Resources <0.01 - 77

CDN-GS-P7B CDN Resources 0.71 0.07 304

CDN-GS-P7E CDN Resources 0.766 0.086 38

CDN-GS-2J CDN Resources 2.36 0.2 17

CDN-GS-3G CDN Resources 2.59 0.18 197

CDN-GS-5F CDN Resources 5.3 0.36 395

CDN-GS-1OC CDN Resources 9.71 0.65 2

CDN-GS-11A CDN Resources 11.21 0.87 61

S A M P L E S E C U R I T Y 11.3

11.3.1 Sample Security Prior to 2009

SRK has no information regarding the sample security during exploration work prior to 2009.

11.3.2 Sample Security: 2009 to Present

Guyana Goldfields maintains a well-documented chain of custody for the assay samples submitted for

assaying. Drill core is under the control of drilling contractors who deliver core boxes to the Aurora camp

for logging. The logging and sampling areas are secured by a fence. Once samples have been taken, they

are securely packed for shipment in sealed rice bags. During shipment, which is carried out by company

personnel in company-owned vehicles, sample batches are accompanied by sample submission forms.





Table 11–2: Specific Gravity Database for the Aurora Gold Project

Domain

Saprolite* Fresh^

Number ofSamples

Mean

SpecificGravity

StandardDeviation

Number

ofSamples

Mean

SpecificGravity

StandardDeviation

Aleck Hill 16 2.78 0.06 Aleck Hill North 30 2.85 0.07Rory's Knoll 93 2.81 0.06Walcott Hill 29 2.80 0.11Walcott Hill East 17 2.80 0.06Mad Kiss 17 2.76 0.06Mad Kiss South 9 2.76 0.08Mad Kiss West 25 2.74 0.20

Weighted Average 49 1.73 0.23 236 2.80* Based on new AMEC saprolite data inclusive of quartz vein material. Sample locations unknown.

^ No update in database since 2008.





1 2 . 0 D ATA V E R I F I C A T I O N

V E R I F I C A T I O N B Y G U Y A N A G O L D F I E L D S 12.1

Guyana Goldfields and their independent consultants completed several verification programs for the

preparation of previous technical reports including Cargill and Gow (2003) and Cargill (2005),

Mukhopadhyay (2007) and Montejo et al (2009).

Montejo et al (2009) reviewed the analytical quality control data acquired between 2004 and 2008. Thereview included independent auditing of the exploration database and the performance of assaying

results delivered by the primary laboratories used by Guyana Goldfields. During the review, AMEC

identified problems in the assay results delivered by Loring and forced the implementation of several

improvements, including re-assaying of a large percentage of the samples originally assayed by Loring,

the use of commercial exploration database software and enhancements to the sample handling

practices to avoid mix-ups and mislabeling. In late 2009, Guyana Goldfields retained AMC Mining

Consultants (Canada) Ltd (AMC) to review quality control issues and the presence of coarse gold in

particular.

In July 2009, Guyana Goldfields made two changes to enhance the integrity of the project data;

switching the primary assay laboratory from the non-accredited Loring laboratory to the accredited

Acme and implementing a Century Systems master database, which incorporates stringent data security

protocols.

Since late 2009, Guyana Goldfields has implemented all the recommendations expressed by




NI 43 101 Technical ReportUpdated Feasibility Study | Aurora Gold Project |Guyana, South America

12.2.2 Database

SRK was provided with a GEMS database containing updated borehole data produced during the period

May 2011 to April 2012. SRK was also provided with revised gold mineralization wireframes for all the

gold mineralized domains except for Rory’s Knoll in DXF exchange and GEMS formats. SRK worked with

Guyana Goldfields to define criteria for the definition of these revised gold mineralization domain

wireframes.

SRK conducted a series of routine verifications to ensure the reliability of the electronic data provided by

Guyana Goldfields. These verifications include checking the borehole data for minimum and maximum

values for each field and confirming/editing those outside of the expected ranges; checking for

inconsistency in lithological unit terminology and/or gaps in the lithological code, and checking for gaps,overlaps and out of sequence intervals for both assays and lithology tables.

For the wireframes crossovers, duplicate triangles, gaps and edge boundary joining were verified. SRK

found the GEMS database to be in good order and well maintained. SRK considers the database suitable

for resource estimation. After review, SRK considers that the gold mineralization wireframes interpreted

by Guyana Goldfields represent adequate boundaries for the gold mineralization and can be used as

resource domains for this study.

12.2.3 Verification of Analytical Quality Control Data

SRK analyzed the analytical quality control data accumulated by Guyana Goldfields for the period from

May 1, 2011 to May 31, 2012.

Guyana Goldfields provided SRK with external analytical control data containing the assay results for the

quality control samples for the Aurora Gold Project. All data was provided in Microsoft Excel

spreadsheets.





Table 12–1: Summary of Analytical Quality Control Data Produced by Guyana Goldfields on the Aurora Gold

Project between November 30, 2010 and July 31, 2011

Total (%) Comment

Sample Count 27,688Blanks 1,017 3.67%

Blank 827 Coarse gravelCDN-BL-4 113 CDN Resources (<0.01 ppm Au)CDN-BL-7 77 CDN Resources (<0.01 ppm Au)

Standards 1,014 3.66%CDN-GS-P7B 304 CDN Resources (0.71 ppm Au)CDN-GS-P7E 38 CDN Resources (0.766 ppm Au)

CDN-GS-2J 17 CDN Resources (2.36 ppm Au)CDN-GS-3G 197 CDN Resources (2.59 ppm Au)

CDN-GS-5F 395 CDN Resources (5.3 ppm Au)CDN-GS-10C 2 CDN Resources (9.71 ppm Au)CDN-GS-11A 61 CDN Resources (11.21 ppm Au)

Field Duplicates 988 3.57% Quarter coreTotal QC Samples 3,019 10.90%

Check Assays 0.00%

A number of field blank samples did not return values below detection limit at Acme Labs; assuming a

threshold limit of five times the detection limit less than percent of blanks failed. There are a number of

blank standard samples above the recommend value of less than 0.01 ppm; approximately 10 % of CDN-

BL-4 and 30 % of CDN-BL-7 failed. These blank failures cannot be explained by mislabeled reference

materials. The field blank and standard blank charts (see Appendix B) may present evidence of sample

contamination during the preparation process.

All the gold standard reference materials performed as expected within two standard deviations. There

were only two failures, (measured as a value exceeding two times the standard deviation of the





1 3 . 0 M I N E R A L P R O C E S S I N G A N D

M E TA L L U R G I C A L T E S T I N G

This section is a review of metallurgical test reports completed for the Aurora Gold Project by Lakefield,

Ontario-based SGS Mineral Services (SGS). Outcomes of the testwork were combined with a new mine

plan and process facility design strategy supplied by Guyana Goldfields to form the basis of this

feasibility study. Additional testwork performed on saprolite ore samples was not used in this feasibility

due to the elimination of saprolite-only processing. However, the recent testwork at RDI with Saprolite

ores indicated no problems with rheology up to 55% solids and the material tends to settle reasonablywell.

M E T A L L U R G I C A L P R O C E S S 13.1

The Aurora Gold Project is being developed by Guyana Goldfields with feed from open pit and

underground mining operations. The gold ores will be processed starting at a nominal rate of 5,000 tpd,

expanding to 10,000 tpd at 92% availability. The process facility design feed grade is 3.3 g/t Au, with an

average LoM grade of 2.7 g/t Au. A saprolitic material caps the Fresh Rock orebody in varying thickness

throughout the resource; however the plan is to mine and process both materials simultaneously.

The process flowsheet includes three stage crushing, followed by grinding in a ball milling circuit

operating in a closed-circuit with cyclones. The cyclone overflow slurry will be fed to a pre-leach

thickener and then into a CIL circuit.

The precious metal will leach in the cyanide solution and it will be adsorbed onto the carbon media. The




3 0 ec ca epo tUpdated Feasibility Study | Aurora Gold Project |Guyana, South America

The programs included mineralogy, comminution studies, head assays, gravity separation, leach

extraction, rheology and thickening, and cyanide detoxification testwork. Table 13 –1 lists the reports

completed to date and used for the feasibility study.

Table 13–1: List of Metallurgical Reports

Laboratory Report NameIssueDate

SGS Lakefield Minerals Services,Lakefield, Ontario

An investigation into the Recovery of GoldProject 11198-001 – Report 1

28-Mar-06

AMEC Report


A pre-feasibility investigation into The Solid-LiquidSeparation and Rheology of the Aurora ProjectProject 12088-002 – Final Report

10-Sep-09


An investigation into The Characterisation of Samplesfrom the Aurora ProjectProject 12088-002 – Final Report

14-Sep-09

SGS Lakefield Minerals Services,

Lakefield, Ontario

An investigation into The Recovery of Gold from the Aurora Project Samples

Project 12088-001 Final Report

24-Nov-09

Contract Support ServicesRed Bluff, California

Drop Weight Test Report on Three Samples from Aurora Apr-10

Comminution Dynamics Lab-McGill UniversityMontreal, Quebec

Aurora Project Grinding Media Wear 5-May-10


An investigation into Grindability Testing of Samples fromthe Aurora ProjectP j t 12088 005 G i d bilit R t

26-Jul-10




pUpdated Feasibility Study | Aurora Gold Project |Guyana, South America

Table 13–2: List of SRK Litho-tectonic Domains in Relation to Guyana Goldfields Logs

SRK Domains Guyana Goldfields Lithologies

Interbedded Metasedimentary and Mafic

Metavolcanic Rocks

Ash Tuff, Metasediment, Volcanics/Volcanics Sediments,

Metavolcanics

Mafic Metavolcanic Rocks Metavolcanics, Volcanics/Volcanics Sediments

Metasedimentary Rocks Ash Tuff, Metasediment, Felsic Tuff

Quartz and Feldspar Porphyry Dike Quartz Feldspar Porphyry, Andesite Porphyry, Diorite

Shear Zones Sericite Schist, Chlorite Schist, Felsic Tuff

Tonalite Tonalite, Quartz Vein

13.2.1 Sample Selection

A total of six HQ boreholes were requested by AMEC in 2009 for the April 9, 2012 feasibility study. The

location and the details of the drill holes are presented in Table 13 –3 and Figure 13-1.

Particular focus was made on material likely to be mined in the first five years. The location, azimuth,

and dip of the boreholes were considered to maximize the number of variability composite samples that

could be produced and to also provide an even spatial representation of the different areas of the

deposit.

Table 13–3: Boreholes Co-ordinates

BoreholeDrill Hole Co-ordinates Relative Level

Northing Easting (m)MET-AH-1 750963.494 195915.656 103.621MET-AH-2 750870.212 195912.516 98.115MET-EW-1 751540.384 196623.556 55.353MET-RK-1 751556.557 196796.738 54.604MET MK 1 751128 106 196412 323 61 397





Source: Ausenco, 2012

Figure 13–1: Boreholes Locations and Details





Table 13–4: Head Analysis

CompositeSampleNumber

Calc. HeadGrade Au, g/t

Agg/t

ST %

S=

%

Rory’s Knoll Saprolite 1 0.47 - 0.02 < 0.05Rory’s Knoll Saprolite 1 - 0.02 - 0.14 < 0.05Rory’s Knoll Upper Volcanics 2 0.26 - 0.18 0.11Rory’s Knoll Upper Volcanics 1 - 0.73 < 0.5 0.10 < 0.05Rory’s Knoll Sericite Chlorite Schist 3 2.01 - 0.39 0.23Rory’s Knoll Tonalite 4 0.08 - 0.16 0.12Rory’s Knoll Quartz Vein 5 2.11 - 0.67 0.56Rory’s Knoll Lower Volcanics 6 0.27 - 0.48 0.26Mad Kiss Upper Volcanics 7 0.02 - 0.05 < 0.05Mad Kiss Quartz Feldspar Porphyry 8 3.35 < 0.5 0.79 0.58Mad Kiss Quartz Felsic Tuff 9 0.03 - 0.06 < 0.05Mad Kiss Quartz Vein 10 1.34 - 0.54 0.34Mad Kiss Diorite 11 1.51 - 0.47 0.31 Aleck Hill Saprolite 12 0.67 < 0.5 0.38 < 0.05 Aleck Hill Saprolite 2 - 2.13 < 0.5 0.03 < 0.05 Aleck Hill Saprolite 3 - 0.06 - 0.02 < 0.05 Aleck Hill Upper Volcanics 13 12.4 - 1.15 0.99 Aleck Hill Lower Volcanics 14 0.68 - 0.26 0.17 Aleck Hill Quartz Vein 15 0.30 - 0.12 0.10

13.2.3 Sample Preparation

Four sample shipments were received by SGS between January 2006 and March 2010, for a total gross

weight of approximately 3,755 kg. Five samples were prepared by SGS for McGill University

Comminution Dynamics Lab. The sample preparation of the shipments received by SGS is described

herein. SGS project 12088-005, shipment received on March 2010:

A total of 162 drill core samples from Rory’s Knoll, Aleck Hill and Mad Kiss deposit (~ 3 tons);




Updated Feasibility Study | Aurora Gold Project |Guyana, South America

Six samples received in two pails (35kg).

TE S T W O R K R E V I E W 13.3

Tetra Tech reviewed the results from various test programs to verify the process design criteria and

flowsheets designed by Ausenco which were used as the basis for this report. Section 17 discusses

process design details.

13.3.1 Mineralogy

In general, the gold mineralization occurs in the vein systems, whose characteristic mineralization and

lithology for the four zones (in the wireframe model) is described in Table 13 –5 (updated from previous

interpretations as part of the work conducted by SRK in 2011).

Quartz-ankerite veins containing trace pyrite associated with hydrothermal alteration as

chief occurrence; and

Auriferous veins from weak to moderate stockwork preferentially in competent lithologies.

Table 13–5: Lithology and Mineralization by Ore Zone, Aurora Gold Project (Mineral Resource Evaluation,

SRK)

Deposit Lithology and Mineralization

Rory’s Knoll

Auriferous veins in highly altered porphyritic diorite, intensely carbonated, albite altered(tonalite).

Abundant gold-rich quartz-ankerite veins in tonalite near sheared zone on westerncontact of tonalite with barren mafic schists.

Vein density increases towards contact.

W l tt Hill E t if t b t i i ili i h it t fi t t





Table 13–6: Crystalline Mineral Assemblage Phases of the S

aprolite and

Fresh Rock Composite

Samples (12088-001)

Saprolite Composite Sample Fresh Rock Composite SampleMajor Quartz Quartz

Moderate Mica, plagioclase PlagioclaseMinor Kaolinite, potassium feldspar Dolomite, mica, siderite, potassium feldsparTrace Hematite, goethite Calcite, chlorite, pyrite

13.3.2 Comminution Characteristics

Three main series of comminution testing were undertaken to determine the characteristics of theAurora Gold Project deposit.

The first series of comminution tests were conducted at SGS as part of project 12088-001. The Saprolite

and Fresh Rock “master” composites were submitted for a basic Bond ball mill work index test (at 100

mesh, 150 µm). The results are shown in Table 13 –7.

Table 13–7: Bond Ball Mill Grindability Test Summary (12088-001)

Sample Name Mesh ofGrind

F80 P80 Gram perRevolution

BWi HardnessPercentile(mm) (mm) (kWh/t)

Saprolite Comp 100 1269 78 3.20 7.0 1.1Fresh Rock

Comp100 2129 116 1.69 14.2

47.0

The Saprolite composite indicates a low hardness, while the Fresh Rock composite indicates a moderate

degree of hardness.





done on individual lithology samples. No testing was done on the saprolite sample from Rory’s Knoll and

Aleck Hill. Test results from the SGS report are shown in Table 13-9.

The Rory’s Knoll and Mad Kiss composites had similar hardness characteristics, with indices falling in themedium to hard categories, while the Aleck Hill ore was slightly softer. The bond abrasion indices

generally fell in the medium range of the SGS database. The MacPherson steady-state mill charge built

up some coarse and rounded pebbles in the mill, but nothing critical to restrain throughput rate.

The third testwork series was performed at the McGill University Comminution Dynamics Lab in order to

predict the wear of the grinding media. The results are presented in Table 13-10.

The testwork results showed that for both the ball milling and the previous SAG milling configurationsused in the analysis, the quartz vein ore caused the greatest overall wear compared to the other ore

types. Increasing the top size from 25 mm to 50 mm of the media in the ball mill decrease the predicted

media wear by more than 40%.

Table 13–9: Comminution Test Summary (12088-005)

Sample NameRelativeDensity

JK ParametersMacPherson

TestAWi CWI RWi BWi (kWh/t) Ai

Axb Axb ta (kg/h) (kWh/t) (kWh/t) (kWh/t) (kWh/t) 170M 200M (g)

Rory’s Knoll Comp 2.81 31.2 35.6 0.38 9 8.8 14.8 15.5 - 14.6 - -S1-RK Saprolite - - - - - - - - - - - -S2-RK Upper Volcanics 2.74 - 37.1 0.35 - - - - 16.1 12.9 - -S3-RK Sericite Chlorite Sc. 2.86 - 36.5 0.33 - - - - - 13.6 - 0.16S4-RK Tonalite 2.84 - 41.9 0.38 - - - - 15.8 12.5 13.2 0.148S5-RK Quartz Vein 2.76 - 39.8 0.37 - - - - 14.7 13.9 13.7 0.348S6-RK Lower Volcanics 2.78 - 36.1 0.34 - - - - - 12.6 - -Mad Kiss Comp 2.85 32.3 32.3 0.39 8.6 9.3 14.6 16.8 - 14.5 - -S7-MK Upper Volcanics - - - - - - - - - 15.2 - -S8-MK Quartz Felspar Por. 2.78 - 46.3 0.43 - - - - - 16.7 - -S9-MK Quartz Felsic Tuff 2.81 - 43.9 0.4 - - - - - 12.3 - -S10 MK Q t V i 14 5



NI 43-101 Technical ReportU d t d F ibilit St d | A G ld P j t |G S th A i




The third series of tests were done by SGS as part of project 12088-005. A series of gravity separation

tests using a Knelson MD-3 concentrator, followed by Mozley mineral separator, were conducted to

produce tailings for downstream cyanidation testwork. The results are presented in Table 13-13.

Table 13–13: Gravity Separation Test Results (12088-005)

TestNo.

Sample NameHead Calc.

(Au g/t)Feed Size

K80 (μm) Recovery

(% Au)Tailings(Au g/t)

Concentrate(Au g/t)

G130% Aleck Hill Saprolite 2/70%Rory’s Knoll Quartz Vein Blend

1.53 138 32.0 1.04 246

G230% Aleck Hill Saprolite 2/70%Rory’s Knoll Upper Volcanics 1Blend

0.69 135 24.3 0.52 274

G3 30% Aleck Hill Saprolite 2/70%Rory’s Knoll Sericite ChloriteSchist Blend

1.60 132 28.8 1.14 320

G630% Aleck Hill Saprolite 2/70%Mad Kiss Diorite Blend

1.59 ~150 24.7 1.20 301

G4 Aleck Hill Saprolite 2 1.98 ~150 8.4 1.82 47.3G5 Rory’s Knoll Upper Volcanics 1 0.49 ~150 38.7 0.30 250

G7Mad Kiss Quartz FeldsparPorphyry

2.33 ~150 70.4 0.69 2113

The recovery of gold by gravity separation ranged from 8% for Aleck Hill saprolite 2% to 70% for MadKiss quartz feldspar porphyry.

Another series of gravity separation test using only a Knelson MD-3 concentrator were conducted to

produce gravity concentrate for downstream intensive cyanidation testwork. The results are presented

in Table 13-14.

The recovery of gold from the Knelson concentrate ranged from 18% to 54%.

NI 43-101 Technical ReportU d t d F ibilit St d | A G ld P j t |G S th A i




13.3.4 Leaching

Intensive Cyanidation

Intensive cyanidation testwork was conducted by SGS for project 12088-005. The tests were conductedon gravity concentrates, (Table 13-14) and the results are summarized in Table 13-15. Gold extraction by

intensive cyanidation ranged from 88% to 98% after 48 hours of leaching. Cyanide consumption was

high, most likely due to the high peroxide additions required to maintain the oxygen levels above 20

ppm.

Table 13 –15: Intensive Cyanidation Results Summary

Test No. Sample Name

NaCN Cons.

(kg/t)

% Au Recovery Residue

(Au g/t)8h 24h 48hILR-1 Rory’s Knoll Upper Volcanics 321 88 89 88.0 4.01ILR-2 Rory’s Knoll Quartz Vein 161 93 94 93.4 3.46

ILR-3 Rory’s Knoll Saprolite 146 86 80 90.0 0.14

ILR-3R Rory’s Knoll Saprolite Sample 1 14 95 97 98.0 0.40

ILR-4 Rory’s Knoll Saprolite 2 48 88 93 96.5 0.96

ILR-530% Aleck Hill Saprolite 2/70% Rory’s Knoll Quartz Vein Blend

111 87 88 88.6 11.0

Whole Ore Cyanidation

Whole ore cyanidation tests were completed by SGS for project 12088-001. The tests were performed

on the Saprolite and Fresh Rock composites and each sample was tested at three different grind sizes.

Results are summarized in Table 13-16.

Table 13–16: Whole Ore Cyanidation Results Summary (12088-001)

CN Feed Reagent Cons Extraction

NI 43-101 Technical ReportUpdated Feasibilit St d | A rora Gold Project |G ana So th America




Table 13–17: Bulk Whole Ore Cyanidation Test Results (12088-005)

SampleTestNo.

K80 CN Feed

calcReagent Cons.

(kg/t)Extraction

(% Au)

(μm) (Au g/t) NaCN CaO 48hFresh Rock Composite CN1 74 4.20 0.42 0.42 94.5Fresh Rock Composite CN2 74 4.39 0.77 0.32 91.1Fresh Rock – Golder Paste Tech Sample CN4 58 0.41 0.89 0.23 92.7Fresh Rock – Golder Paste Tech Sample CN5 58 0.75 0.90 0.29 96.0Fresh Rock – Golder Paste Tech Sample CN6 58 0.69 0.97 0.31 97.1Saprolite Composite CN3 58 3.51 0.53 1.83 97.6

Leach Gravity Tailings

Three series of leach tests were conducted to determine the effect of the grind size, the retention time,

sodium cyanide (NaCN) concentration and the effectiveness of lead nitrate (Pb(NO3)2) addition.

The first series of leach tests were conducted by SGS for project 12088-001. Six tests were completed on

the two composite samples to determine the effect of grind size and retention time on gold recoveries.

The testwork was completed on gravity tailings produced from the gravity separation testwork. The

results of the leach tests are summarized in Table 13-18.

For the Saprolite composite, the gold recovery ranged from 90% to 97% after 48 hours of leaching. For

the Fresh Rock composite, the gold recovery ranged from 85% to 94% after 48 hours of leaching. The

gold recoveries increase with the fineness of the grind.

Table 13–18: Summary of Cyanidation Test Results on the Saprolite and Fresh Rock Composites Gravity

Tailings (12088-001)

CNRecovery Reagent Cons Extraction

NI 43-101 Technical ReportUpdated Feasibility Study | Aurora Gold Project |Guyana South America




Table 13–19: Gravity Tailings Cyanidation Results (12088-002)

Test

No.

K80 CN Feed

Calc.

Recovery,Gravity

Conc.

RetentionTime

NaCNConc.

Reagent Cons.(kg/t)

Extraction Residue

(μm) (Au g/t) (% Au) (h) (g/L) NaCN CaO (% Au) (Au g/t)CN1 55 2.25 50.4 48 0.5 0.23 0.49 91.6 0.19CN2 76 2.3 50.4 48 0.5 0.08 0.44 89 0.25CN3 82 2.42 50.4 8 0.75 0.27 0.21 85.1 0.36CN4 82 1.98 50.4 24 0.75 0.18 0.31 90.9 0.18CN5 82 1.94 50.4 32 0.75 0.25 0.33 90.2 0.19CN6 81 2.12 50.4 48 0.75 0.26 0.36 90.8 0.2CN7 72 2.19 50.4 24 0.5 0.28 0.3 90.4 0.21CN4 82 1.98 50.4 24 0.75 0.18 0.31 90.9 0.18CN8 72 2.4 50.4 24 1 0.35 0.23 91.9 0.2CN9 72 2.02 50.4 24 1.5 0.47 0.22 90.3 0.2

The third series of leach test were conducted by SGS for project 12088-005. Twelve tests were

completed on blended samples to determine the effect of the grind size, lead nitrate addition and NaCN

concentration. The tests were conducted on gravity tailings. The results of the leach test are presented

in Table 13-20 and Table 13-21.

Recovery for gold ranged from 93% to 97%. The addition of lead nitrate and the fineness of the grind didnot appear to have any significant impact on the gold recovery. Reducing the NaCN concentration to 0.2

g/L reduced the gold recovery to 65%.

Table 13–20: Cyanidation of Gravity Tailings Test Results – Effect of Grind Size and Lead Nitrate Addition

(12088-005)

Sample No.CN Test

No.K80 Pb(NO3)2 Reagent Cons. (kg/t) Recovery

% AuResidue

Au, g/t(μm) (kg/t) NaCN CaO





Table 13–22: Cyanidation of Gravity Tailings – Variability Testing (12088-005)

Test No.CN Test

No.K80 NaCN Conc. Reagent Cons. (kg/t)

Recovery% Au

Residue

(μm) (g/L) NaCN CaO 24h 36h 48h Au, g/tG5 CN20 61 0.75 0.16 0.88 88 83 88.3 0.04G7 CN22 53 0.75 0.38 0.57 94 94 92.7 0.05G6 CN21 76 0.75 0.09 1.44 88 88 90 0.12

Leaching and Adsorption Kinetics

Leaching and gold cyanide adsorption kinetic tests were conducted by SGS for project 12088-002 to

investigate the effect of pulp density. A leaching kinetic test was performed on gravity tailing from the

master composite. The adsorption kinetic test was performed on a pulp generated from a bulk cyanide

leach. Results of the test are presented in Table 13-23, Table 13-24, Figure 13-2 and Figure 13-3. Results

from the tests indicated that pulp density (in the range of 45% to 55% solids) has no effect on either the

kinetics of leaching or the gold cyanide adsorption.

Table 13

–

23: Leach Kinetic Results Summary (12088-002)

Pulp Density(% solids)

Leach Time(h)

Extraction Au(%)

Solution ResidueAu, g/t(Au, mg/L)

45

2 48.3 0.99 1.3104 67.0 1.38 0.8108 80.6 1.67 0.490

24 92.7 1.93 0.18548 92.3 1.93 0.196

50

2 50.8 1.30 1.2804 67.7 1.74 0.8408 80.4 2.08 0.510





Table 13–24: Adsorption Kinetic Results Summary (12088-002)

Pulp Density(% solids)

Leach Time(h)

Solution LoadingAu, g/t(Au, mg/L)

45

0 1.93 00.5 1.73 2291 1.60 3762 1.41 5864 1.09 9347 0.78 1265

10 0.58 147424 0.20 186548 0.10 1969

50

0 2.42 00.5 2.02 372

1 1.92 4682 1.62 7304 1.25 10527 0.86 1384

10 0.62 158324 0.21 191548 0.10 2006

55

0 2.97 00.5 2.21 564

1 2.08 658

2 1.86 8124 1.29 12017 0.88 1472

10 0.63 163324 0.19 190748 0.10 1965





Source: Ausenco, 2012

Figure 13–2: Leach Kinetic as a Function of Pulp Density Plot

40

50

60

70

80

90

100

0 10 20 30 40 50 60

% A u E x t r a c t i o n

Retention time, h

45% solids 50% solids 55% solids

2000

2500





Table 13–25: Settling Thickening Test Results Summary (12088-002)

SampleFlocculentCIBA

Dosage Feed U/F(% wt)

TUFUA(m

2 /t/day)

THUA(m

2 /t/day)

ISR(m

3 /m

2 /day)

SupernatantVisual(g/t) (% wt)

MasterComp Magnafloc 455 10 10 67 0.07 0.01 566 Clear

Saprolite Comp

Magnafloc 10 50 10 52 0.04 0.01 1639 Clear

Table 13–26: Rheology Test Results Summary (12088-002)

Sample Test Solids %Unsheared Sample

ηP Sheared Sample

ηP γ τyB γ τyB

(Range, s-

) (Pa) (mPa.s) (Range, s-

) (Pa) (mPa.s)

Master Comp

T1 71.8 480-600 124.0 96 300-600 38.9 78T2 69.2 360-600 34.2 85 360-600 11.6 75T3 66.4 480-600 15.4 35 200-400 5.9 35T4 62.4 300-600 3.8 21 120-300 2.2 20

Saprolite Comp-20mesh

T5 61.7 200-240 124.0 13 360-600 63.3 49T6 57.0 400-600 49.8 50 240-480 27.8 37T7 54.1 360-600 19.9 20 240-400 13.2 26T8 50.4 360-600 7.5 15 240-480 4.2 18T9 46.0 120-400 2.2 9 120-360 1.7 8

The data indicated that the critical solid density for the master composite (thickener underflow) was in

the range of 68% to 69% wt, at a corresponding yield stress of ~34 Pa. The critical solids density for the

Saprolite composite -20 mesh was in the range of 54-55% wt, at a corresponding yield stress of ~20 Pa.

Additional rheology and settling thickening response testwork was conducted on three different

samples. The results summary is presented in Table 13-27 and Table 13-28. The maximum underflow

solids densities predicted based on rheologically determined critical solid density were 47% wt for the





Table 13–27: Settling Thickening Test Results Summary (12088-005)

Aleck Hill,Saprolite 3,

-10 mesh

Blend*Rory’s

Knoll

Sampling Test 24 16 25 8Particle Size, d80, μm 29 31 31 48

Pulp pH 6.5 7.4 10.5 7.2

Initial Solids, % wt 5.8 10 10 10.0

CIBA Magnafloc 10, g/t dry 77 103 50 45

U/F Solid Density, % wt 40 48 53 53

*Max. U/F % wt Predicted by CSD 47 50 nd 60

Thickener U/F Unit Area, m /t/day 0.06 0.08 0.16 0.06

**Above corrected for CSD actual UF 0.07 0.08 nd 0.07

Thickener Hydraulic Unit Area, m /t/day 0.01 0.01 0.03 0.01**Above corrected for CSD actual UF 0.01 0.01 nd 0.01

Initial Settling Rate, m3/m

2/day 1649 833 265 1567

Supernatant Clarity, 10’-60’-final Clear Clear Clear Clear

* 30% Aleck Hill Saprolite 3/70% Rory’s Knoll Upper volcanics

Table 13–28: Rheology Test Results Summary (12088-005)

Sample Test Solids %Unsheared Sample

ηP Sheared Sample

ηP γ τyB γ τyB

(range, s-

) (Pa) (mPa.s) (Range, s-

) (Pa) (mPa.s)

Rory’s Knoll pH ~ 7.2

T1 63.3 200-250 131 Plug 200-400 51 63T2 60.5 200-400 42 42 200-400 27 41T3 56.3 200-400 14 28 200-400 11 23T4 50.8 200-400 4 15 200-400 3 13

Aleck HillT5 50.7 200-400 94 58 200-400 73 56T6 47.8 200-400 43 45 200-400 43 34





In continuous testing on 250-360 mg/L CNWAD solutions, 4.1 g to 5.0 g SO2/g CNWAD and 2.0 g to 2.5 g

hydrated lime/g CNWAD, with minor addition of copper (from copper sulphate), achieved less than 2 mg/L

CNWAD.

13.3.7 Metallurgical Recoveries

The metallurgical recoveries used for both reserve estimations and financial analysis are based on all

completed testworks. The recoveries used were 97.0% for the Saprolite and 94.7% for the Fresh Rock.

The results represent an average for all the Aurora Gold Project deposits.

13.3.8 Conclusions

The conclusions and recommendations drawn from the testwork program are presented below:

Tonalite ore in Rory’s Knoll and volcanics at Aleck Hill represent the major mineralized rock

types in run of mine (RoM) ore feed. Other rock types represent minor components. Values

for design should be weighted according to the major ore types;

Ores are highly amenable to cyanide leaching with recovery percentages in the low to mid-

90s (excluding those samples where the grinds were too coarse). Finer grinds of P 80 of 75

µm, enhanced leach recovery. Cyanide consumption from the laboratory tests averaged 0.5

kg/t for fresh rock and is typical for a free milling ore with few deleterious cyanide

consumers;

Leaching and carbon adsorption kinetic tests indicated there should be no effect of

increased pulp density (in the range of 45% to 55% solids) on either parameter; and

The ore is amenable to detoxification by the air/SO2, copper catalyzed process (air/SO2/Cu2+)

with industry-normal reagent demand and acceptable CNWAD levels for the discharge to a

wet tailings dam.





1 4 . 0 M I N E R A L R E S O U R C E E S T I M AT E S

M I N E R A L R E S O U R C E E S T I M A T I O N M E T H O D O L O G Y 14.1

The evaluation of mineral resources for the Aurora Gold Project involved the following procedures:

Database compilation and verification;

Resource modeling:

Updating wireframe model for Rory’s Knoll by SRK and importing of 3D wireframe models

for the other deposits received from Guyana Goldfields; Extensive validation of database and the wireframe models prepared by Guyana Goldfields;

Data processing (compositing and capping) and statistical analysis;

Selection of estimation strategy and estimation parameters;

Block modeling and grade interpolation;

Validation, classification and tabulation;

Assessment of “reasonable prospects for economic extraction” and selection of reporting

cut-off grades; and Preparation of Mineral Resource Statement.

D A T A B A S E 14.2

14.2.1 General

Data used to evaluate the mineral resource were provided by Guyana Goldfields as a Microsoft Excel

and drawing exchange (DXF) files




p y y | j | y ,

The GEMS database was found to be in good order and well maintained. On completion of the validation

procedure, SRK considers the database suitable for resource estimation.

The wireframes were validated by checking for crossovers, duplicate triangles, gaps in the wireframesand edge boundary joining. SRK accepted wireframe definitions and resource domains defined by

Guyana Goldfields for all zones other than Rory’s Knoll which was modeled by SRK.

R E S O U R C E M O D E L I N G P R O C E D U R E S 14.3

14.3.1 Geological Model

The Aurora Gold Project is subdivided into nine distinct auriferous zones: Aleck Hill, North Aleck Hill,

Rory’s Knoll, East Rory’s Knoll, Walcott Hill, East Walcott Hill, Mad Kiss, South Mad Kiss, and West MadKiss.

The nine distinct zones are grouped into four main auriferous zones shown in Figure 14-1. The gold

mineralization in Rory’s Knoll (and to some extent in East Rory’s Knoll and East Walcott) form “carrot”

shaped outlines containing a weak to moderate stockwork of quartz-carbonate veins. The gold

mineralization at Aleck Hill and Mad Kiss areas forms distinct tabular zones of quartz-carbonate veins.




p y y | j | y ,

Figure 14–1: Oblique Section Looking North Showing the Main Auriferous Zones of the Aurora Gold Project

The Rory’s Knoll domain was modeled by SRK, considering logged lithology and gold grade distribution

patterns. In addition to re-defining the outline of this domain, SRK also remodeled zones of internalwaste located inside the Rory’s Knoll domain. In this process, several previously modeled internal waste

zones were removed except where demonstrated continuous based on geological data. Those internal

waste zones were not considered for grade estimation.

The Rory’s Knoll domain was also subdivided into a high grade subdomain by constructing wireframes

inside the main wireframe around areas of higher grade mineralization. The limits of the higher grade

subdomain were initially modeled using Leapfrog using a threshold of 5.0 g/t gold. The Leapfrog meshes

were subsequently manually smoothed to define more consistent zones of higher grade goldmineralization. The resulting Rory’s Knoll high grade subdomains are entirely contained within the main

Rory’s Knoll domain.

SRK has additionally defined the top 10 m layer of saprolite generated using the Leapfrog shells based

on the 0.2 g/t gold cut-off. The Guyana modeled wireframes representing the vein like structures were

subdivided into the fresh and saprolite domain using the bottom surfaced of the saprolite provided to

SRK by Guyana.

Structural geology investigations, geological modeling, and information from the infill drilling completed

during the period of May 2011 to April 2012 support a better definition of the lateral continuity of the

structures hosting the gold mineralization within all the gold zones, and consequently improve the

confidence in the geological continuity of the gold mineralization.

Infill drilling completed on the other auriferous zones (Aleck Hill, Mad Kiss, and East Walcott Hill, the

“satellite deposits”) prompted revision to the geological interpretation. The boundaries of the gold




p y y | j | y ,

Mad Kiss Saprolite 722South Mad Kiss 800South Mad Kiss Saprolite 822West Mad Kiss 900

West Mad Kiss Saprolite 922Saprolite Horizontal (10 m) 2222

14.3.2 Database Preparation

The domain wireframes were used to code a zone field into the block model (Table 14-1). The geological

solids were coded and these values were written into the block model using the wireframe to delineate

the auriferous zones. Table 14-2 illustrates the basic sample gold grade and sample length statistics for

the borehole data.

Unsampled borehole intervals intersecting geological wireframes were assigned a value of 0.003 g/t.

Metallurgical and geotechnical boreholes were excluded from the database prior to the estimation.

Table 14–2: Basic Statistics of Raw Borehole Samples for the Aurora Gold Project

Domain Variable Count Minimum Maximum Mean Std. Dev. Variance COV Aleck Hill

Au(g/t)

4,187 0.00 122.44 2.82 6.86 47.12 2.44 Aleck Hill North 1,092 0.00 45.77 2.10 4.06 16.47 1.93Rory’s Knoll 14,360 0.00 532.50 2.92 8.35 69.70 2.86Rory’s Knoll HG 3,239 0.00 532.50 5.32 15.34 235.30 2.88Rory’s KnollEast

401 0.00 123.47 4.54 8.37 70.10 1.84

Walcott Hill 534 0.00 80.13 1.89 4.61 21.24 2.44Walcott Hill East 2,249 0.00 2343.93 3.95 49.85 2484.97 12.61Mad Kiss 688 0.00 157.00 4.24 11.28 127.25 2.66Mad Kiss South 218 0.01 41.30 2.46 4.34 18.86 1.77Mad Kiss West 229 0.00 150.20 3.37 11.23 126.06 3.33Saprolite




y y | j | y

Source: SRK, 2012

Figure 14–2: Sample Length Histograms for Aleck Hill, Aleck Hill High Grade, Rory’s Knoll and Rory’s Knoll

High Grade.




Source: SRK, 2012

Figure 14–3: Sample Length Histograms for Aleck Hill North, Rory’s Knoll East, Walcott and Walcott East.




Source: SRK, 2012

Figure 14–4: Sample Length Histograms for Mad Kiss, Mad Kiss West and Mad Kiss South.




For each domain, a capping value was determined by analyzing histograms and cumulative frequency

plots of gold composites (Figure 14-5, Figure 14-6 and Figure 14-7). Capping values were adjusted

iteratively by reference to summary statistics to ensure the robustness of statistics to chosen capping

values.










Basic statistics for the raw and capped gold composites are shown in Table 14-3.

Table 14–3: Statistics for Raw and Capped Gold Composites

Category Nb. Min Max Mean Capped SD VAR COVComp. Raw Cap Raw Cap Raw Cap # cut % Raw Cap Raw Cap Raw Cap

Aleck Hill 4,877 0.00 0.00 83.32 70.00 2.39 2.38 4 0.08% 7.00 6.73 48.99 45.33 2.38 2.3

Aleck Hill North 1,085 0.00 0.00 45.54 30.00 1.91 1.89 3 0.28% 23.31 19.64 543.56 385.89 0.48 0.42

Rory's Knoll 13,637 0.00 0.00 412.84 80.00 2.76 2.72 7 0.05% 4.26 4.11 18.18 16.86 1.79 1.74

Rory’s Knoll HG 2,874 0.00 0.00 412.84 80.00 5.11 4.90 6 0.05% 6.74 4.83 45.41 23.36 2.22 1.62

Rory's Knoll East 434 0.00 0.00 67.49 30.00 3.45 3.36 3 0.69% 11.07 6.77 122.64 45.78 2.17 1.38

Walcott Hill 535 0.00 0.00 49.60 15.00 1.62 1.50 4 0.75% 5.65 5.08 31.9 25.77 1.61 1.47

Walcott Hill East 2,259 0.00 0.00 542.18 40.00 2.60 2.33 4 0.18% 3.42 2.07 11.73 4.27 2.13 1.4Mad Kiss 702 0.00 0.00 138.57 60.00 3.89 3.68 4 0.57% 15.86 4.06 251.39 16.46 4.9 1.42

Mad Kiss South 228 0.02 0.02 20.69 15.00 2.19 2.13 3 1.32% 9.22 7.47 85.02 55.83 2.41 2.04

Mad Kiss West 272 0.00 0.00 150.20 20.00 3.07 2.22 6 2.21% 2.56 2.2 6.58 4.83 1.28 1.14

Saprolite Hor. 2,197 0.00 0.00 54.90 20.00 0.73 0.71 4 0.18% 4.29 3.17 18.43 10.08 1.37 1.09

14.3.4 Specific Gravity

Specific gravity was measured by Guyana Goldfields on 465 pieces of fresh core using a waterdisplacement methodology (Table 14-4). An additional forty-nine measurements from saprolite were

generated by AMEC. Specific gravity does not vary much between various auriferous zones. An average

specific gravity of 2.80 was assigned to all fresh waste rock blocks, and 1.73 for all saprolite blocks above

the saprolite wireframe. SRK recommends that more specific gravity measurements be generated for

each zone as the above data was acquired from only a few drill holes.

Table 14–4: Guyana Goldfields Specific Gravity Measurements




Principal directions were initially determined by the orientation of the data; sensitivities were evaluated

by varying the direction specification and comparing the resulting experimental variograms.

For each domain, SRK examined four different spatial metrics: (1) traditional semi-variogram, (2)traditional correlogram, (3) normal scores semi-variogram, and (4) normal scores correlogram. In

general, the correlogram and normal scores transform facilitate the identification of spatial structure in

the composite data, particularly when the traditional variogram shows little continuity. Variogram

modeling was performed by assessing the structure(s) apparent from these different spatial measures,

and fitting the most reliable measure. Whenever possible, the traditional variogram is the preferred

measure to fit a model; however, the correlogram and/or normal scores variogram are often fitted due

to the noise apparent in the traditional variogram.

The fitted variogram models were cross-checked against the mineral wireframes within Gemcom to

ensure consistency in orientation and reasonableness for estimation purposes. Modeled variograms for

each resource domain are presented in Appendix C. Modeled variogram parameters are tabulated in

Table 14-5.

Table 14–5: Variogram Models for the Aurora Gold Project

Zone C0 C1 C2 C3 R1x R1y R1z Mod1* R2x R2y R2z Mod2* R3x R3y R3z Mod2

Aleck Hill 0.20 0.70 0.10 13.0 13.0 8.5 Exp 65.0 65.0 8.5 SphSaprolite 0.20 0.50 0.30 15.0 15.0 5.0 Sph 40.0 40.0 10.0 Sph

Aleck Hill North 0.20 0.65 0.15 16.0 25.0 8.0 Exp 80.0 25.0 8.0 Sph

Rory’s Knoll 0.25 0.25 0.28 0.22 6.0 8.0 3.0 Exp 6.0 30.0 30.0 Exp 100 30 30 Sph

Rory’s Knoll HG 0.25 0.25 0.50 5.0 5.0 20.0 Sph 34.0 34.0 20.0 Sph

Rory’s Knoll East 0.20 0.60 0.20 70.0 11.0 5.0 Exp 70.0 70.0 5.0 Sph

Walcott Hill 0.20 0.60 0.20 12.0 50.0 6.0 Exp 50.0 50.0 6.0 Sph

Walcott Hill East 0.20 0.33 0.47 5.0 50.0 4.0 Sph 50.0 50.0 4.0 Sph






boundaries are considered hard boundaries for the estimation of the high grade sub-domains, but are

considered a soft boundary for the estimation of the remaining domain (e.g. lower grade envelope

surrounding the high grade sub-domains).




Table 14–8: Grade Estimation Search Parameters

ZoneAleck

HillSaprolite

NorthAleck Hill

Rory'sKnoll

Rory’s Knoll HG

East Rory'sKnoll

Code 200 2222 300 100111

500Pass 1

No. composites (min/max) 3/8 3/8 3/8 3/8 3/8 3/8

Type of search Octant Octant Octant Octant Octant Octant

Minimum number of octants 2 2 2 2 2 2

Max composite per octant 5 5 5 5 5 5

Max composite per borehole 2 2 2 2 2 2

Search radius about X 65 40 80 100 35 70

Search radius about Y 65 40 25 30 35 70

Search radius about Z 12 10 10 30 20 10

Pass 2

No. composites (min/max) 2/12 2/12 2/12 2/12 2/12 2/12

Type of search Ellipse Ellipse Ellipse Ellipse Ellipse Ellipse

Minimum number of octants - - - - - -

Max composite per octant - - - - - -

Max composite per borehole - - - - - -

Search radius about X 130 80 160 200 70 140

Search radius about Y 130 80 50 60 70 140Search radius about Z 24 20 20 60 40 20

Pass 3

No. composites (min/max) 2/20 2/20 2/20 2/20 - 2/20

Type of search Ellipse Ellipse Ellipse Ellipse - Ellipse

Minimum number of octants - - - - - -

Max composite per octant - - - - - -

Max composite per borehole - - - - - -




ZoneWalcott

HillEast Walcott Hill

MadKiss

South MadKiss

West MadKiss

Code 600 400 700 800 900

Max composite per borehole - - - - -

Search radius about X 100 100 180 110 100

Search radius about Y 100 100 70 110 100

Search radius about Z 24 24 20 20 20

Pass 3

No. composites (min/max) 2/20 - - 2/20 2/20

Type of search Ellipse - - Ellipse Ellipse

Minimum number of octants - - - - -

Max composite per octant - - - - -

Max composite per borehole - - - - -Search radius about X 150 - - 165 150

Search radius about Y 150 - - 165 150

Search radius about Z 36 - - 30 30

14.3.8 Resource Model Validation

As a validation check of the ordinary kriging estimates, gold grades were also estimated using an inverse

distance estimator. Results from the two estimators were compared visually and both estimators deliververy similar results. SRK prefers to report gold grades estimated by ordinary kriging because the spatial

continuity and nugget effect can be modeled using variograms and also because ordinary kriging delivers

an estimate of the quality of the estimates in the form of the kriging variance.

The model was further validated visually by comparing block grade estimates to informing capped

composite data on vertical sections and elevation plans. The statistics of the informing capped

composited data also compare well to that of the estimated resource blocks






The mineral resources for the Aurora Gold Project are reported at a cut-off grade of 0.30, 0.40 and 1.80

g/t gold based on open pit (saprolite and fresh rock) and underground mining scenarios, respectively.

The open pit cut-off grades are based on assumptions summarized in Table 14-10, while the

underground reporting cut-off grades was determined considering the same price and recoveryassumptions in consultation with SRK mine engineers involved in the design of an underground mine for

the Aurora Gold Project.

Table 14–10: Conceptual Pit Optimization Assumptions Considered for Open Pit Resource Reporting

Parameter Assumption

Saprolite Fresh

Pit Slopes (per geotechnical sector) 23 to 31 degrees 43 to 51 degrees

Mining cost (ore and waste) US$1.40/t US$1.75/tIncrement mining cost by 5 m bench US$0.02/t mined/bench US$0.02/t mined/bench

Process cost US$6.00/t feed US$8.00/t feed

G & A costs US$3.00/t feed US$4.00/t feed

Process recovery 97.0 percent 94.7 percent

Assumed process rate 4,000 tpd 8,000 tpd

Gold price US$1,300 per ounce US$1,300 per ounce

Mining dilution 16.0 percent 16.0 percent










M I N E R A L R E S O U R C E S T A T E M E N T 14.4

Mineral resources were classified according to the CIM Definition Standards for Mineral Resources and

Mineral Reserves (December 2005) by Dorota El-Rassi, PEng (APEO #100012348), Dominic Chartier,

PGeo (Ordre ded Geologue du Quebec #874), and Glen Cole, PGeo (APGO#1416), appropriate

independent qualified persons for the purpose of NI 43-101.

The Mineral Resource Statement for the Aurora Gold Project is summarized in Table 14-11. Mineral

resources are reported at two cut-off grades to reflect the fact that parts of the gold mineralization are

amenable for open pit extraction, while other parts are more likely amenable for underground

extraction. The mineral resources for the Aurora Gold Project are reported at a cut-off grade of 0.30,

0.40 and 1.80 g/t gold based on open pit (Saprolite and Fresh Rock) and underground mining scenarios,

respectively. In Table 14-11 the nine gold deposits are grouped for reporting as follows:

Rory’s Knoll - includes Rory’s Knoll, Rory’s Knoll East and Walcott Hill East;

Walcott Hill - includes Walcott Hill;

Mad Kiss - includes Mad Kiss, South Mad Kiss and West Mad Kiss; and

Aleck Hill - includes Aleck Hill and North Aleck Hill.




Table 14–11: Mineral Resource Statement*, Aurora Gold Project, Guyana, SRK Consulting (Canada) Inc.,

June 25, 2012

Classification Zone


000’ Tonnes Au (g/t) 000’ Ounces

Sap Fresh Total Sap Fresh Total Sap Fresh Total

Open Pit Mining

MeasuredRory’s Knoll 0.15 5.62 5.77 3.30 3.23 3.23 0.02 0.58 0.60

Total Measured 0.15 5.62 5.77 3.30 3.23 3.23 0.02 0.58 0.60

IndicatedSaprolite 2.73 0.00 2.73 1.15 0 1.15 0.10 0 0.10 Aleck Hill 2.10 11.60 13.69 2.54 2.57 2.56 0.17 0.96 1.13

Rory’s Knoll 0.36 7.74 8.10 2.87 2.64 2.65 0.03 0.66 0.69Walcott Hill 0.05 0.62 0.67 2.37 2.05 2.08 0.00 0.04 0.04

Mad Kiss 0.26 1.56 1.81 2.03 3.61 3.38 0.02 0.18 0.20Total Indicated 5.49 21.51 27.01 1.84 2.66 2.49 0.33 1.84 2.16

Total M + I 5.64 27.13 32.77 1.88 2.77 2.62 0.34 2.42 2.76

InferredSaprolite 2.35 0.00 2.35 0.93 0 0.93 0.07 0 0.07 Aleck Hill 0.28 1.34 1.61 1.55 1.67 1.65 0.01 0.07 0.08

Rory’s Knoll 0.03 0.33 0.36 2.31 2.60 2.57 0.02 0.03 0.05

Walcott Hill 0.01 0.06 0.07 2.78 1.72 1.89 0.00 0.00 0.00Mad Kiss 0.21 0.51 0.72 2.21 3.00 2.77 0.00 0.05 0.05

Total Inferred 2.89 2.23 5.12 1.10 2.11 1.54 0.10 0.15 0.25

Underground Mining

Indicated

Aleck Hill 0 2.70 2.70 0 3.83 3.83 0 0.33 0.33

Rory’s Knoll 0 25.68 25.68 0 3.89 3.89 0 3.21 3.21

Walcott Hill 0 0.18 0.18 0 2.80 2.80 0 0.02 0.02Mad Kiss 0 1.50 1.50 0 4.60 4.60 0 0.22 0.22

Total Indicated 0 30.06 30.06 0 3.91 3.91 0 3.78 3.78

Total M + I 0 30.06 30.06 0 3.91 3.91 0 3.78 3.78

Inferred Aleck Hill 0 1.62 1.62 0 3.78 3.78 0 0.20 0.20




The impact of the 2012 drilling program can be assessed by comparing the September, 2011 and June

2012, resource statements. This sensitivity is tabulated in Table 14-12.

Table 14

–

12: Impact of the 2012 Drilling Program on the Resource Statement

2011 2012 Difference

ClassificationQuantity Grade

ContainedAu

Quantity GradeContained

AuQuantity Grade

ContainedAu

000’ Tonnes Au g/t000’

Ounces000’ Tonnes Au g/t

000’Ounces

000’Tonnes

Au g/t000’

Ounces

Open Pit Mining

Measured 5.75 3.29 0.61 5.77 3.23 0.60 0.29% -1.80% -1.81%Indicated 14.47 3.31 1.57 27.01 2.49 2.16 86.64% -24.79% 37.69%Inferred 3.48 3.41 0.39 5.12 1.54 0.25 47.08% -54.79% -34.94%

Underground Mining

Measured 0 0 0 0.00 0.00 0.00 0.00% 0.00% 0.00%Indicated 26.82 4.09 3.52 30.06 3.91 3.78 12.07% -4.36% 7.39%

Inferred 6.49 3.74 0.78 11.81 4.12 1.56 82.03% 10.06% 100.46%

Combined MiningMeasured 5.75 3.29 0.61 5.77 3.23 0.60 0.29% -1.80% -1.63%Indicated 41.29 3.83 5.10 57.06 3.24 5.94 38.20% -15.44% 16.51%Inferred 9.97 3.63 1.17 16.93 3.34 1.82 69.83% -8.05% 55.32%

The mineral resources are highly sensitive to reporting cut-off grade. To illustrate this sensitivity, the

block model quantities and grade estimates are presented at various cut-off grades in Table 14-13. The

reader is cautioned that these figures should not be misconstrued as a mineral resource. The reported

quantities and grades are only presented as a sensitivity of the resource model to the selection of cut-off

grade. This cut-off grade sensitivity is also illustrated as grade tonnage curves shown in Figure 14-10.




Table 14–13: Global Block Model Quantity and Grades* Estimates at Various Cut-off Grades, Aurora Gold

Project, Guyana

Cut-offMeasured & Indicated Inferred

Quantity Grade Quantity Grade(g/t gold) ('000 tonnes) Gold (g/t) ('000 tonnes) Gold (g/t)0.10 83.26 2.69 25.91 2.600.20 82.76 2.70 25.88 2.610.30 81.89 2.73 25.81 2.610.40 80.64 2.77 25.65 2.630.50 78.91 2.82 25.34 2.650.60 76.90 2.88 24.95 2.690.70 74.56 2.95 24.42 2.730.80 72.01 3.02 23.70 2.790.90 69.31 3.11 22.81 2.87

1.00 66.63 3.20 21.71 2.961.20 61.19 3.38 19.26 3.201.40 55.87 3.58 16.55 3.511.60 50.85 3.79 14.12 3.861.80 46.11 4.00 12.73 4.092.00 41.73 4.22 11.56 4.313.00 24.66 5.44 7.16 5.514.00 15.14 6.70 4.47 6.715.00 9.55 8.01 2.96 7.906.00 6.31 9.32 2.30 8.60

7.00 4.35 10.61 1.61 9.498.00 3.11 11.86 1.21 10.189.00 2.25 13.15 0.90 10.7410.0 1.67 14.43 0.64 11.25

* The reader is cautioned that the figures presented in this table should not bemisconstrued as a Mineral Resource Statement. The reported quantities and grades areonly presented as a sensitivity of the deposit model to the selection of cut-off grade.




1 5 . 0 M I N E R A L R E S E R V E E S T I M AT E S

I N T R O D U C T I O N 15.1

The open pit and underground ore reserves only include measured and indicated resources. The

methodologies and modifying factors utilized to generate these reserves are described in sections 16.2

and 16.3 for open pit and underground respectively.

M I N E R A L R E S E R V E S T A T E M E N T 15.2

The SRK mineral reserve estimate is shown in the table below. It is based on the June 2012 resource

estimate prepared by SRK. This reserve estimate includes the results from open pit mining in Rory’s

Knoll and the surrounding satellite pits, as well as underground mining in Rory’s Knoll.

Table 15–1: Mineral Reserve Statement*, Aurora Gold Project, Guyana, SRK Consulting (Canada) Inc.,

January 11, 2013

Quantity Grade Contained AuProven (kt) (g/t) (k oz)

OP SAP 168 2.64 14

OP FRESH 2,207 3.07 218

Total Proven 2,375 3.04 232

Probable

OP SAP 4,955 1.70 270

OP FRESH 6,343 3.03 618

Undergro nd 25 851 2 84 2 357




1 6 . 0 M I N I N G M E T H O D S

H Y D R O G E O L O G Y 16.1

A numerical, three-dimensional (3D) finite-element groundwater flow model was prepared by Itasca

Denver, Inc., (Itasca) to predict inflow to the open pits and the underground mine workings at Aurora

mine. The model was constructed based on available data from various site investigations.

Five open pits and one underground sublevel retreat (SLR) mine were simulated for the Feasibility Study.

Six shear zones were located in the model. These shear zones potentially intercept either open pits or

the SLR mine. Both the open pits and the SLR are located a few hundred meters south of the Cuyuni

River. Figure 16-1 shows the model domain, locations of the mine layout, footprint of the shear zones,

the Cuyuni River, and monitoring points. At this stage, there are no field-observed hydraulic conductivity

values obtained for the shear zones.

There are three major geologic units in the model domain. These geologic units include the

unconsolidated deposits, weathered bedrock, and fresh bedrock. The thickness of the unconsolidated

deposits ranges from a few meters to approximately 50 meters (m). The thickness of the weatheredbedrock ranges from a few meters to approximately 20 m and was simulated in the model with a

constant thickness of 5 m. The fresh bedrock is less permeable than the weathered bedrock and the

unconsolidated deposits.

Man-made dikes (Figure 16-1) were simulated in the model. These dikes are assumed to extend from

the ground surface to the top of the fresh bedrock and were simulated as low-permeability materials.




Figure 16

–

1: Base Map for Guyana Site




Model boundary conditions were simulated as follows:

A recharge range of approximately 7 to 9 mm/year (based on the model calibration) was

applied to the water table.

The river was simulated as a constant head boundary condition.

Regional groundwater flow through lateral boundaries was simulated as a variable flux

boundary condition.

The bottom of the model was simulated as a no-flow boundary condition.

A steady-state model calibration was conducted to match the measured water levels in both the

unconsolidated deposits and the bedrock. The calibrated hydraulic parameters are summarized in Table

16-1. The calibrated hydraulic conductivity (K ) values are within the range of the measured values asshown in Figures 16-2, 16-3, and 16-4.

Table 16–1: Hydraulic Parameters of Geologic Units in Groundwater Flow Model

Formation/Unit

Hydraulic Conductivity

(m/day)Specific

Storage

(m-1

)

Specific

Yield

( )K x K y K z

Unconsolidated

Deposits

< 1400 m from the riverbank 1.0E-01 1.0E-01 1.0E-02 1.0E-05 2.0E-01

> 1400 m from the riverbank 1.8E-01 1.8E-01 1.8E-02 1.0E-05 2.0E-01

Weathered

Bedrock

< 200 m from the riverbank 5.0E+01 5.0E+00 5.0E+00 1.0E-06 1.0E-02

between 200 and 400 m from the riverbank 4.0E-01 4.0E-02 4.0E-02 1.0E-06 1.0E-02


> 600 m from the riverbank 1 0E 02 1 0E 03 1 0E 03 1 0E 06 1 0E 02




Figure 16–2: Measured Kh and Modeled Kx in Unconsolidated Deposits Based on Distance from River




Figure 16–4: Geometric Mean of Kh and Modeled Kx vs. Depth

The open pits and the SLR were simulated with a series of drain nodes that were activated according to

the excavation schedule of each pit. The configurations of the ultimate pits are shown in Figure 16-5.

The mining schedules for these pits are summarized in Table 16-2.




Table 16–2: Open Pits Excavation Schedule

Pit NameBottom Elevation

(mamsl)Excavation Schedule(Production Years)

Start End

Rory's Knoll

SAP 25 1 3

PB1 -35 4 5

PB2 -95 4 6

Aleck Hill

SAP 30 1 4

PB1 -35 4 7

PB2 -110 4 14

Aleck Hill NorthSAP 45 2 3

PB1 0 8 13

Mad KissSAP 40 4 4

PB1 -60 11 15

Walcott Hill PB1 35 14 15

The ramp was assumed to start in Year 3 and the SLR was assumed to start in Year 5. The zone of

relaxation (ZOR) related to SLR was provided by SRK, Inc. (SRK) based on its geomechanical model

simulation. The development of the ZOR over the life of mine (LOM) was simulated in the groundwater

flow model by increasing the K value of the rock within the ZOR. The simulated K value for ZOR is

summarized in Table 16-1.




hydraulic parameter values for the base case scenario are summarized in Table 16-1. Given the limited

knowledge available from site-specific testing on the anisotropic ratio and shear zones, Itasca simulated

three additional scenarios to demonstrate the degree of sensitivity and the possible range of inflows

associated with different parameter values for the man-made dike, the shear zone, and the anisotropicratio.

The predicted inflow rates for the base case scenario according to the different open pits are shown in

Figure 16-6. The predicted maximum inflow rate is approximately 1,600 and 600 m3/day to the Aleck Hill

Pit and Rory's Knoll Pit, respectively. The predicted inflow rate to the SLR underground workings is

shown in Figure 16-7. The predicted inflow rate increases from 1,700 m3/day in Year 6 to 2,000 m3/day

by the end of mining.




Figure 16–7: Base Case Scenario: Simulated Inflow Rates to SLR Workings

The predicted inflow rates for the three sensitivity simulations are shown in Figures 16-8, 16-9, and 16-

10 and are summarized by Itasca (2013).




Figure 16–9: Sensitivity Analysis: Inflow to Aleck Hill Pit




As mentioned above, surface runoff was not simulated in the model. Management of surface runoff is

discussed in a later chapter.

OP E N P I T M I N I N G 16.216.2.1 Open Pit Mine Geotechnical

A majority of the information, analysis and design parameters provided within the open pit geotechnical

section have been derived from previous work completed by AMEC (2009) and compiled by SRK (2012)

in an earlier 43-101 document, released in early 2012. Design parameters for the hard rock portions of

the proposed pits were reviewed and considered to be robust enough to allow for the proposed changes

made to the open pit design (i.e. shallower pit depths); therefore, additional analyses were not

conducted for this updated version of the previous report. The following section summarizes thegeotechnical information relevant to open pit mining.

The geomechanical site investigation program to aid in the feasibility level mine design of the open pits

for the Aurora Gold Project consisted of:

Twelve NQ vertical geotechnical diamond drill holes totaling 1,146 m, which were not

oriented but geotechnically logged through the overburden and geomechanically logged for

approximately 50 m into the fresh bedrock; Eleven inclined geomechanical open pit NQ holes totaling 2,602 m of drilling, of which all

core was oriented and geomechanically logged;

Six main inclined geomechanical underground NQ holes (2 holes were wedged from existing

geomechanical holes) totaling 3,550 m of drilling, of which all core was oriented and

geomechanically logged;

A total of 29 exploration holes intersecting the tonalite orebody at Rory's Knoll, totaling




sericite schist unit (BH10-RK-RMU-05 and BH10-RK-RMU-06) were also packer tested to identify

hydrogeological parameters for the sericite schist.

The data from these boreholes was used to determine material strength, rock mass classification [NGI-Q

classification (Barton et. al., 1974), CSIR Rock Mass Rating, RMR'76 and RMR'89, (Bieniawski, 1976,

1989)] and the Geological Strength Index, GSI (Hoek et. al., 1995).

Note the pit modeled for the Updated Preliminary Economic Assessment (AMEC, 2009) was the primary

pit geometry used for determination of kinematic, planar and wedge stability analysis and specification

of design criteria where originally based on this. Confirmation of the final pit wall stability based on the

updated pit geometry for the feasibility level pit was reviewed and found to be satisfactory based on

these original design criteria. The following sections present summaries of the key material properties

used for design of the open pits, some of which may also be relevant to the design of the underground

mining areas.

Overview of Geotechnical Domains

Based on geological interpretation available during the original analyses for the open pit and

underground mining areas (Figure 16-11), eight main structural domains were identified. The first two

domains are both within the volcanic sediments but were differentiated based on the structural

features; volcanic sediments with moderate foliation and the massive volcanic sediments that have lessprominent foliation and random calcite veining. These domains comprise the majority of the country

rock:

Domain 1: Volcanic sediments (VCS). This is the predominant rock type, but has been found to also form

inter-bedded layers within the massive volcanic sediments. This unit is found within both the Aleck Hill

and Rory’s Knoll areas.




thickness varying from 5 m to 50 m can be identified to the south-west of the tonalite pipe, and extends

from surface down to a depth of 1400 m, but appears to have limited lateral extent.

Domain 6: Ash Tuff (AST). This rock unit is identified in Rory’s Knoll and it is mainly present in the upper

parts of the orebody and appears to form a halo around the tonalite pipe. It is characterized by more

significant dark ash bands interbedded with Felsic Tuff, and has strong to moderate foliation present.

Domain 7: Felsic Tuff (FLS). Again this unit is primarily found at Rory's Knoll and appears to form a halo

around the pipe. As with the Ash Tuff, foliation is moderate to strong and greater than in the volcanics

sediments. The main difference between this unit and the Ash Tuff is the less significant ash banding.

Both the Ash Tuff and the Felsic Tuff tend to be slightly stronger than the volcanic sediments.

Domain 8: Diorite (DIO). Minor diorite dike intrusions were identified in exploration and geomechanicalholes, having a thickness of 5 m to 20 m. These dikes are assumed to strike east-west in the general

trend of the foliation based on the jointing, but have not been found to bisect the tonalite pipe;

suggesting their formation was prior to the intrusion of the pipe. These intrusive dikes of Diorite and

Granodiorite are more prevalent at Aleck Hill.





b h l l d ll d d f h k l ll ( ) l f ld



boreholes were also drilled to identify the rock mass properties at Walcott Hill (WH). Only field point-

load tests were performed at this location. For the other three potentials, small pits of Aleck Hill North,

Mad Kiss and Mad Kiss South, the properties of the nearest rock units have been assumed for this

study.., Representative rock core specimens of all rock units were collected and sent to AMEC’sHamilton Laboratory for uniaxial compressive strength – UCS (89 tests), density (108 measurement,

generally 8 per rock type), Young’s modulus and Poisson’s ratio (22 generally 4 per rock type), Brazilian

tensile strength (55), triaxial strength testing (30 tests, 6 tests per rock type and 2 tests per

confinement), multi-stage direct shear testing (6 tests) and point-load testing (35 tests, 5 per rock type).

The procedures for all rock testing followed ISRM suggested methods and ASTM corresponding standard

procedures. Additionally, during the core logging campaign in the field, a total of 985 point-load tests

were performed. The results of these tests are summarized in Table 16-3.

Table 16–3: Summary for the Material Testing of Each Rock Type (AMEC, 2012).


d i i d i T bl 16 4 f R ' K ll Sli h i i h id ifi d



domains are summarized in Table 16-4 for Rory's Knoll. Slight variation on these sets was identified at

Aleck Hill with a clockwise rotation of approximately 15°.

Table 16

–

4:

Summary of Major Joint Sets per Domain of Rory’s Knoll AMEC, 2012).

North-south jointing was identified throughout the project location from all of the 2,700 joints oriented

in the Rory's Knoll area, 400 in Aleck Hill and 340 in Walcott Hill. This suggests that most of the structure




ith f th h i l t di f d t il d d i It i t d th t d i th d l t f



with further geomechanical studies for detailed design. It is expected that during the development of

any pit, continued geomechanical mapping and verification of design versus actual performance is

necessary.

Based on the RMR values, the rock mass at Aurora is of a good to very good quality. The high rock mass

competency the main rock pit slopes are predominantly controlled by structural jointing and not by the

material strength. The exception to this is the weathered overburden, consisting of residual soils,

saprolites and saprolitic rock (saprock), in which the material strength is the main controller of the

stability of slopes. The residual soils and saprolites can typically vary in thickness from 5 m to

approximately40 m around the pit limits, with maximum thicknesses of up to around 80 m is found at

the centre of Aleck Hill.

The objective of the pit wall stability analysis is to determine, through kinematic analysis, optimum

bench face angles (BFA). Additionally, consideration is given to planar, wedge and toppling failure, with

selection of the bench width to obtain an appropriate inter ramp angle (IRA), and achieve an effective

catch bench width. Overall slope stability, based on the overall slope angle (OSA), has been performed

using probabilistic limit equilibrium techniques, considering the rock mass strength and hydrogeological

conditions. The definition of BFA, bench width and height, IRA, and OSA used here are summarized in

Figure 16-12.


most sensiti e pit alls (see Fig re 16 11) The stabilit anal sis carried o t for each section considered



most sensitive pit walls (see Figure 16-11). The stability analysis carried out for each section considered

three main cases; 1.) completely unsaturated (well drained or dry) conditions, 2.) using phreatic surface

based on a hydrogeological model, and 3.) For completely saturated (undrained) conditions. This latter

condition is a worst case scenario and assumes a constant recharge from surface water and no drainage.

Given the potential for drainage to the underground, the most realistic case is based on the

determination of the phreatic surface (Table 16-6). This was based on simplified hydrogeological

permeability models develop from field packer testing results. Rory's Knoll bedrock permeability ranges

from k = 5 x 10-6 to 1 x 10-7 m/s reducing with depth, while lower permeability was found on the

western side of the property at Aleck Hill with bedrock permeability ranging from k = 3 x 10-6 to 5 x 10-9

m/s.




Table 16–6: Summary of the High Wall SLIDE Analyses for Segment 1, 4 and 5 (AMEC, 2012).

The probabilistic analysis was based on the variability of the rock properties: UCS, GSI, and unit weight.

The Generalized Hoek-Brown strength criterion was employed to determine the failure plane shear

strength. Additionally, two disturbance factors (Hoek et. al., 1995) were considered; D = 0, which

represents fresh intact rock as the upper bound case and D = 0.7, which represents degradation of the

rock mass by large scale production blasting as the lower bound case. For this study, a seismic

acceleration was considered using a value of 0.1 g.

Typically, the minimum design factors of safety (FoS) for non-critical pit slopes is 1.2; however, here due

to the high importance of maintaining stability of the north wall of Rory's Knoll, as a failure could result

in destabilization of the river dike, a higher tolerance of 1.3 was applied to this zone. Only the most

sensitive slope stability sections, Section 1, Section 4 and Section 5, are summarized in Table 16-7.

A ramp segment at Rory's Knoll’s north wall (Segment 1, Figure 16-14) reduced the OSA to 48.3°. This


Bedrock 3 70 15 6 52 6 40 6 to 42 9 ---



Bedrock 3 70 15 6 52.6 40.6 to 42.9 ---

*OSA dependant on number of ramp segments that pass through the wall. The main ramp is assumed to be27m and pit bottom ramp 20m. OSA is also measured to the bedrock contact.

*Zone 5 could be increased to 34˚ for slopes <40m high if upper bound strengths are confirmed bylaboratory testing.

In the north wall of Rory's Knoll (Design Zone 2), toppling failure will predominate, and a maximum

practical BFA of 75° has been recommended. It is assumed that toppling cannot be avoided for most

practical bench face angles as joints are steep (85° to 90°), such that toppling failures will be managed

with an increase in the bench width and regular bench maintenance. Bench widths have been

recommended as 8 m for this zone. This same design criterion has been applied to the north east corner

of Aleck Hill North and the northern sector of Aleck Hill (Design Zone 2).

In the south wall of Rory's Knoll and Walcott Hill (Design Zone 3), it was identified that the dominant

foliation set dips at a shallower angle (70° to 72° versus the pervasive 85° to 90°) and the BFA has been

selected to match the mean foliation dip at 70° with a minimum bench width of 6 m. For Aleck Hill, it

was identified that plane wedge failure on sub vertical joints under saturated conditions and the

development of a tension gash could occur in both the north east sector and the south west sector

(Design Zone 2's Aleck Hill). These potential failures may result in small bench scale wedges, if tension

cracks developed within 1.5 m to 0.5 m of the crest, respectively of the zone. The bench width has beenincreased to 8 m to accommodate potential failure; however, with drained conditions these should be

stable.

For all other regions, based on the identified jointing, no other kinematically permissible failures were

identified and, for these zones (Design Zone 1), a practical BFA of 75° has been recommended with a

minimum bench width of 6 m.


Analyses were performed using (C=20 kPa Φ=37°) for material strength A bench face angle (BFA) based



Analyses were performed using (C=20 kPa, Φ=37 ) for material strength. A bench face angle (BFA) based

on site experience of 70° is recommended, which was found to be stable at site for excavations 5 m to 7

m high.

Saprolite Pit Slope Design Recommendations

Based on a limit equilibrium slope stability analysis, the open pit design criteria used for this study is

summarized in Table 16-9 below. These zones correlate to the design sectors shown in Figure 16-16.

NIC3

4




Figure 16 –15 Fresh Rock Sectors




Figure 16 –16: Saprolite Sectors


Table 16–8: Recommended Slope Geometry



Rock SlopeSector

BenchFaceAngle

BenchHeight

ApproximateBench Width

OSA/IRAFully

Drained

OSA/IRAPartiallyDrained

(˚) (m) (m) (˚) (˚) Saprolite RK1 65 7 8 34 30

Saprolite RK2 70 8 8 40 36




Saprolite AH1 70 10 8 46 NA

Saprolite AH2 70 10 8 46 NA

Saprolite AH3 70 10 8 46 NASaprolite AH4 70 10 8 46 NA

Saprolite AHN1 70 8 8 39 35




Saprolite MK1 70 8 8 36 32

Additional Design Criteria

It is recommended that for the saprock/weathered bedrock interface a 10 m wide catch bench should

be left.

For the first 15 m bench in weathered rock, an 8 m catch bench is recommended, regardless of the zone.

This is intended to contain any potential loose rock, as it has been observed that there is a general

tendency for the rock mass to have a lower RQD (average 70%) in this slightly weathered zone from 2 m

t 20 thi k d di th i


Bedrock surface weathering has created a saprolite zone over the deposit Below this layer the



Bedrock surface weathering has created a saprolite zone over the deposit. Below this layer, the

mineralization is contained in fresh bedrock. Open pit mining of near surface saprolite and fresh bedrock

mineralization is planned.

16.2.3 Open Pit Optimization

Pit optimization was conducted using Whittle™ software. The software utilizes the Lerchs-Grossmann

algorithm to generate a pit shell that provides the maximum operating margin, or cashflow (before

capital, taxes or discounting), based on a resource model and a set of economic and technical input

parameters.

Pit optimization economic parameters include unit mining costs, processing costs, general and

administrative costs, and unit revenue estimates. Pit optimization technical parameters include pitfootprint constraint, estimates of mining dilution, mining loss, process recovery, and pit overall slope

angles. Pit overall slope angles are derived from geotechnical criteria adjusted for the expected haulage

ramp layout.

In accordance with the guidelines of National Instrument 43-101 Standards of Disclosure for Mineral

Projects and the CIM’s Definition Standards for Mineral Resources and Mineral Reserves, only those ore

blocks classified in the Measured and Indicated categories are allowed to drive the pit optimization for a

feasibility level study. Inferred resource blocks, regardless of grade and recovery, bear no economicvalue and are treated as waste.

A series of nested pit shells was generated by varying or factoring unit revenue estimates (referred to as

revenue factor or RF). Nested pit shells are utilized for incremental and present value analysis, and to

guide phase pit and ultimate pit design.

The Rory’s Knoll pit area was optimized as an open pit/underground mine cross-over. This methodology


Table 16–9: Aurora Pit Optimization Input Parameters



Parameter Saprolite Fresh Rock

Revenue

Revenue basis Measured and Indicated only

Au Price ($/oz) 1,300

Au Payable (%) 100%

Refining Charges ($/oz) 3.00

Royalties (%) 8%

Value of Au in Dore ($/oz) 1193

Value of Au in Dore ($/g) 38.4

Mill Recovery (%) 92% 93%

Value of Au in Plant Feed ($/g) 35.29 35.68

Estimated Operating Costs

Mining Cost at Surface ($/t) 2.62 2.77

Incremental Mining Cost ($/t/bench) 0.04 0.04

Underground Mining Cost ($/t) 22.04

Process Cost ($/t) 5.36 12.77

G&A Cost ($/t) 4.1 3.2

Processing Rates

Plant Feed Rate (Mt/a) 1.75

Discount Rate (%) 8%

Mine Design Parameters

Mining Recovery (%) 95%


Table 16–10: Overall Slope Angles used in the Pit Optimization

BenchTwoWay Total



Domain/ZoneSectorName

BenchFace

Angle(º)

BenchHeight

(m)

BermWidth

(m)InterrampAngle (º)

Wallheight

(m)

# TwoWay

Ramps

WayRampWidth

(m)

# SingleLane

Ramps

SingleLane Ramp

Width (m)

TotalWidth of

Ramps(m)

OverallAngle(º)

Used

Fresh Rock

1 75 20 8 56.3 200 3 17 1 13 64 48

6 75 20 8 56.3 200 1 17 1 13 30 512 75 20 11 50.7 200 1 17 1 13 30 43

7 75 20 11 50.7 200 1 17 1 13 30 46

3 70 20 8 52.6 200 4 17 1 13 81 45

Saprolite

A 65 5 6.5 29.5 30 1 17 0 13 17 26

B 65 5 6.5 30.0 30 1 17 0 13 17 26

C 70 5 6 32.6 30 1 17 0 13 17 28

D 70 5 6 32.6 30 1 17 0 13 17 28

E 70 5 6.5 31.0 30 1 17 0 13 17 27

F 70 5 5 36.2 30 1 17 0 13 17 30

NI 43-101 Technical ReportUpdated Feasibility Study Aurora Gold Project Guyana South America



Updated Feasibility Study Aurora Gold Project Guyana, South America.

Whittle was used to generate a preliminary discounted cashflow based on the economic parameters

presented. Note that the resultant cashflows and NPV generated by the Whittle optimization excludeany initial and sustaining capital cost requirements. The two discounted cashflows are referred to as the

‘Best Case’ and the ‘Worst Case’. The ‘best case’ cashflow assumes that each nested shell will be mined

in sequence. This method will release the highest grade and lowest strip ratio ore first, which results in

the highest possible cashflow; however, mining each shell in sequence is impractical because of narrow

mining widths and maintaining access ramps to all working areas. The ‘worst case’ cashflow assumes

that the whole pit will be mined out bench by bench. This method brings waste stripping forward in time

and produces a lower discounted cashflow. In reality, the mine will be designed in phases which are

practical for the mining equipment selected and the true net present value will likely fall between the

best and worst cases. In order to get a more practical representation of the true net present value, an

average between the best and worst cases is also calculated.

The net present value and quantity of ore and waste for each nested shell are compared in order to

select the best pit shell as the ultimate pit. This analysis was conducted separately for both the Rory’s

Knoll pit as well as all of the satellite pits combined. The graphic comparison (Figure 16-17 below) of the

nested pit shells and the pit shells selected as the ultimate. In both of the following figures, pit shell 36

represents the breakeven pit shell which is generated by revenue factor 1.





In the Rory’s Knoll cross over analysis, pit 20 was selected as the ultimate shell because it contains less

waste than the breakeven pit shell without losing significant tonnes of ore.





Figure 16–19: Aurora-Whittle Pit Optimization Incremental NPV





16.2.4 Open Pit Design

Mine design criteria was based on conventional open pit truck and shovel mining using 7.7 m3 front end

loaders combined with 43.5 tonne articulated haul trucks. The same fleet of loaders and trucks is

planned for both ore and waste.

Pit shell 20 from the Rory’s Knoll underground cross over optimization and pit shell 21 from the standard

pit optimization of the other pits were used to guide the detailed design of the ultimate pits in each

area.

The detailed pit designs incorporate the defined pit slope geotechnical parameters (bench face angle,inter ramp angle and berm width) for the various rock types and pit sectors. The design also accounts for

haul ramp access to all mining areas and minimum practical mining widths based on the selected mining

equipment. The pit design parameters are summarized below.

Table 16–12: Pit Design Parameters

Parameter Unit Value

Two-way Ramp Width m 17Single Lane Ramp Width m 13

Maximum Grade % 10

Minimum Pushback

Width m 40

Bench Height m 5

Saprolite Berm Interval m 5

NI 43-101 Technical ReportUpdated Feasibility Study Aurora Gold Project Guyana, South America.




engineering should consider including these small satellite pits. These excluded zones are in the Mad

Kiss South mineralized zone.

The first phase of each pit was designed targeting the saprolite ore in order to mine and be able to

process this material before mining of the fresh rock. The various saprolite pit designs are shown in

Figure 16-20 below.




p y y j y ,

Figure 16–21: Ultimate Whittle Pit Designs

Aleck Hill is the largest pit area and consists of two fresh rock phases. Rory’s Knoll is the second largest

pit and it is also mined as two fresh rock phases Aleck Hill North and Mad Kiss are smaller pits and they




p y y j y ,

The in-pit sumps and ditches that are required for the open pit dewatering plan have not been designed

in the ultimate phase designs. The dewatering sumps have been sized and located. This modification willneed to be included in further detailed engineering of the project.

16.2.5 Waste Storage Areas

Waste material from the open pits is primarily saprolite and fresh rock with small quantities of surface

organic soil. The waste management plan is to stockpile the waste saprolite and fresh rock as close as

possible to the open pits in order to minimize mining costs and equipment requirements. The organic

soils that require stripping will be stockpiled in designated storage areas to be used in later reclamation

activities.

Waste rock facility (“WRF”) design parameters include:

Haulage ramps are 17 m wide, maximum grade of 10%;

Overall dump slope angle 2.5H:1V or 22 degrees;

Saprolite bank density of 1.73 t/m3 and placed swell factor of 1.15;

Fresh rock bank density of 2.8 t/m3 and placed swell factor of 1.3;

Lift height of 10m.

The designed WRFs have a capacity of 40M m3 and the mine plan requires the storage of 35M m3 of

waste material.

The mine plan also includes hauling waste material to build the dike between Rory’s Knoll and the

Cuyuni River. The proposed mine plan does not haul pit waste material to any other civil works or pit




p y y j y ,

Figure 16–23: Waste Storage Areas

16.2.6 Dilution and Mining Losses

Rory’s Knoll WRF Aleck Hill North WRF

Open Pits

Aleck Hill WRF




p y y j y

Table 16–13: External Dilution, Dilution Grades and Mining Losses for Each Mineralized Zone

Mineralized Zone Mining Losses External Dilution Dilution Grade

(%) (%) (g/t)

Upper Saprolite 5 17 0.10

Saprolite Veins 5 23 0.06

Rory's Knoll Fresh 5 4 0.10

Other Fresh 5 22 0.06

16.2.7 Cut-Off Grades

Cut-off grades were estimated for both saprolite and fresh rock ores and were based on a US$1,300/oz

gold price and the preliminary economic parameters and unit operating costs that are detailed in Table

16-14 below.

Table 16–14: Economic Parameters for CoG Estimate

Parameter Unit Value

Ore Type Saprolite Fresh Rock

Au Price US$/oz 1300

Au Payable % 100

Refining Charges US$/oz 3

Royalties % 8

Value of Au in Dore US$/oz 1193




16.2.8 Open Pit Quantities

The total proven and probable reserves within the open pits are 13.7M tonnes of saprolite and fresh

rock grading 2.55g/t gold. The open pit reserves are reported in Table 16-15 below.

Table 16–15: Open Pit Reserves


Proven (M t) (g/t) (M oz)

OP Sap 0.2 2.64 0.01

OP Fresh 2.2 3.07 0.22

Total Proven 2.4 3.04 0.23

Probable (M t) (g/t) (M oz)

OP Sap 5.0 1.70 0.27

OP Fresh 6.3 3.03 0.62

Total Probable 11.3 2.45 0.89

Total P&P 13.7 2.55 1.12

The proven and probable reserves are reported by pit phase in Table 16-16 below.

Table 16–16: Proven and Probable Reserves


Pit Phase (M t) (g/t) (M oz)

RK Sap 0 8 1 98 0 05






Table 16–18: Open Pit Production Contributions by Area

Ore Tonnes Contained Au Tonnes Mined

Rory's Knoll 39% 41% 20%

Aleck Hill 47% 46% 58%

Aleck Hill North 8% 5% 7%

Mad Kiss 5% 7% 12%

Walcott Hill 2% 1% 3%

16.2.9 Open Pit Production Schedule

The overall open pit and underground production mining schedule was developed in order to feed the

mill 1.75M tonnes of ore per year. The Rory’s Knoll pit is mined first in order to start the underground

mining operation as early as possible. Once Rory’s Knoll is complete, the remaining pits are all mined in

conjunction with the underground operations, to continue feeding the mill 1.75M tonnes of ore per

year. The schedule is also designed to smooth out the annual strip ratio and mining fleet requirements

in order to generate a smooth personnel schedule and minimal new equipment purchases.

Due to the strategy of mining Rory’s Knoll first followed by the all of the other pits, there are two stages

to the mining rate and mining fleet. Rory’s Knoll is the lowest strip ratio pit due to the ore body’s

geometry as well as the fact that it is optimized as an open pit/underground cross over. Therefore, the

first three years of operation have a low strip ratio as well as small equipment requirements. As the

mining schedule excavates all of the other pits, the strip ratio increases along with truck, shovel and drill

requirements


Table 16–19: Open Pit Production Schedule



January 2013 147

2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 Total

SAP Ore Mined (000's dmt) 28 858 - 154 1,804 1,475 463 179 1 160 5,121

Fresh Ore Mined (000's dmt) 894 1,755 1,842 1 346 876 1,186 1,176 477 8,553Total Ore Mined (000's dmt) 28 1,752 1,756 1,996 1,805 1,821 1,542 1,484 1,682 637 13,674

Sap Waste Mined (000's dmt) 69 2,507 90 105 4,388 8,179 6,455 3,101 65 1,460 26,421

Fresh Waste Mined (000's dmt) 3 1,364 4,040 1,973 81 1,997 6,690 9,854 9,071 2,243 37,318

Waste Rehandle (000's dmt) 18 72 72 72 73 73 73 73 73 73 676

Stockpile Reclaim (000's dmt) - - 1 - - - 203 119 505 - 828

Mill Feed Grade (g/t) - 2.33 2.90 2.67 1.59 1.90 3.24 2.17 2.93 4.19 2.55

Milled Tonnes (000's dmt) - 1750 1754 1750 1,750 1,325 1,541 1,484 1,682 637 13,674

Average Recovery (%) 97.0 95.7 94.4 94.6 97.0 96.5 95.2 94.7 94.4 95.1 95.4Total Au

Produced (000's oz) - 126 154 142 87 78 153 98 150 82 1,069




There is pre-production mining in the fourth quarter of 2014 with primary production beginning in the

Rory’s Knoll saprolite in early 2015. The fresh rock phases in Rory’s Knoll will be complete and open pitmining will move into the Aleck Hill, Aleck Hill North and Mad Kiss saprolite pits in 2018. These pits are

all mined concurrently from 2018 to 2023. The Walcott Hill pit is designed as a single phase for both

saprolite and fresh rock and it is mined out entirely in 2023. The pit development sequence is illustrated

in Figure 16-24 below.

Figure 16–24: Walcott Hill Pit, Pit Development Sequence

The saprolite and fresh rock tonnes of ore and waste rock mined by period are illustrated in Figure 16-25

below.

Pit Q4 '14 Q1 '15 Q2 '15 Q3 '15 Q4 '15 Q1 '16 Q2 '16 Q3 '16 Q4 '16 Q1 '17 Q2 '17 Q3 '17 Q4 '17 2018 2019 2020 2021 2022 2023

RK Sap

RK1

RK2

AH Sap

AH1

AH2

AHN Sap

AHN1

MK Sap

MK1

WH




Figure 16

–

26: Typical End of Period Map

RL: 15m

RL: 45m

RL: 45m

RL: -5m

RL: -95m




is doing productive work. Blasting delays, washroom breaks and pre-shift inspections all affect operator

efficiency.

Estimated production equipment mechanical availability (MA), use of availability (UofA) and operator

efficiency (OE) are summarized in Table 16-20.

Table 16–20: Open Pit Mining Equipment, Availability

Equipment Availability and Utilization

MA% UofA% OE%

Drills 90 90 93

Shovels 90 86 90

Truck 85 83 90

Drilling and Blasting

Productivity and cost estimates for drilling and blasting are based on 5m benches. It is recommended

that further detailed engineering examine the opportunity to drill and blast 10m benches in the freshrock which may improve productivity and reduce drilling requirements.

Saprolite material does not require drilling and blasting and therefore it accumulates no drill operating

time or drilling and blasting costs. Saprolite will require definition drilling for grade control (further

discussed in the Grade Control Section of this report)

Drilling productivity is based on rock hardness along with first principles of drill and blast design The




ANFO product in dry areas which will result in a cost savings. It is assumed that a blasting supplier will

provide down hole delivery service as well as bulk explosives trucks and explosives facilities.

Loading

The loading of ore and waste material will be conducted using 7.7m 3 front end loaders. These loading

units have a low capital cost, short lead time and reduced maintenance costs due to having a common

fleet. The 7.7 m3 front end loaders (FEL) are a good productivity match to the 43.5t hauling units.

The loader fleet will start at three units for mining Rory’s Knoll. An additional two units will be required

as the mine expands into the remaining pits.

Hauling

The haulage fleet employed in this plan consists of 43.5t articulated haul trucks. Haul profiles were

calculated as the sum of five components:

bench travel time

ramp travel time

ex-pit travel time

destination spot and dump time

spot and loading time at the loading unit.

Bench travel time is defined for each pit phase and is a function of the average distance from the loading

unit to the access ramp. The ramp travel time is a function of the elevation change from the active

i i b h h l i f h E i l i i l l d f h




Figure 16–27: Open Pit Mine Haul Truck Productivity

Roads and Dumps

The following equipment is required to maintain the haul roads and dumps:

Three Cat D8 230kw track dozers Two additional units in 2018;

-

20

40

60 80

100

120

140

160

180

200

2015 2016 2017 2018 2019 2020 2021 2022 2023

T r u c k P r o

d u c t i v i t y ( t / h r )

Truck Productivity by Period

Truck Productivity




Four 6kw diesel portable light towers.

Grade Control

Grade control in fresh rock for both the Rory’s Knoll tonalite and all the vein hosted mineralization is

anticipated to be based on sampling and visual control in the active dig faces. Sampling of the

mineralized zones will be required to determine whether material in a given vein or tonalite is above or

below the cut-off grade. The upper saprolite zone is not anticipated to be visually controlled in the

active mining face. Grade control definition drilling and sampling will be required to outline zones above

cut-off grade in the upper saprolite.

Grade control drilling will be conducted by a dedicated reverse circulation drill that will drill a 10 m by

10 m pattern to depths of up to 20 m. This will define zones above cut-off and the extra drilling depths

will define several benches of saprolite grade control in a single pass of the drill. Marginally sub-

economic areas of mineralization will also be drilled and sample in order to identify any pockets of

potential ore above cut-off grade.

The technical services team will be responsible for collecting the assay data from definition drilling and

interpreting the results to define the zones above cut-off. Finally, they will indicate those zones in thefield and provide direction to the to the mine operations crew during excavation.

Pit Dewatering

The dewatering infrastructure and equipment is sized to handle ground water inflows and precipitation.

The pit dewatering plan is based on diverting as much surface water as possible away from the open

it th ll ti th t th t d t t th it i dit h d b f




Table 16–22: Open Pit Mine Equipment List

Equipment Model Initial Requirement Expansion Total

Major Production Equipment

7.7 m3 Wheel Loader Cat 988H 3 2 5

43.5t Haul Truck Cat 740B 7 11 18

152mm Rotary DrillCAT MD6240 2 1 3

20,000L Water TruckKenworthL20 1 0 1

Wheel Dozer Cat 824H 1 0 1

Track Dozer Cat D8T 1 2 3*

Motor Grader Cat 14M 2 1 3*

Total 17 17 34

*additional required units are already on site.

Annual equipment requirements for the primary production fleet are illustrated in Figure 16-28 below.




Table 16–23: Open Pit Mine, Personnel Requirements

Open Personnel Year 5 (2019)

Operations Supervision 7

Maintenance

Supervision 5

Technical Services 16

Mine Operations

Drillers 8

Blasters 6

Loader Operators 20

Truck Drivers 68

Equipment Operators 36

Labourers 25

Mine Maintenance

Mechanics 17

Welders 12

Electricians 5Planners 4

Helpers 12

Trainees 4

Total 245

NA Expatriots 2% 4

PP Expatriots 10% 24




16.3.3 Underground Mine Geotechnical

Rory’s Knoll is steeply dipping pipe shaped orebody amenable for both open pit and undergroundmining. The approximate dimensions of the mineralization is 140 m x 100 m in plan with the pipe

extending from surface down to the present mine plan design depth of 1,037 m below ground surface.

Current mine design includes an open pit to an approximate depth of 95 mbsl, and a series of sublevel

retreat (SLR) levels set on 25m intervals.

In 2012, SRK conducted an audit of the existing geotechnical data as part of a technical audit of the

Aurora Gold Project with respect to the SLR mining method. The following are the main points of the

audit:

Ground Water – inflows from the joints and sericite schist region, as well as recharge from

the river would need to be investigated.

Infrastructure Data – additional information would be required north of the deposit to

determine rock conditions for infrastructure.

Shear Zone Properties – additional drilling information would be required in the sericite

shear areas to determine rock mass properties.

Numerical Modeling –numerical modeling analyses would be required to determine

stability of the pipe walls and induced stresses around the SLR infrastructure.

Based on the data collected from the previous geotechnical investigations and the open pit geotechnical

domains, basic geomechanical properties are summarized in Figure 16-29, Figure 16-30, and Figure 16-

31.




Figure 16–29: UCS data ranges for the geomechanical domains.

0.0

50.0

100.0

150.0

200.0

250.0

300.0

U C S ( M P a )

Lithology

Boxplot of UCS Data by Lithology

-S.D

Minimum

Average

Maximum

+S.D




Figure 16–31: Fracture frequency (fractures per metre) data ranges for the geomechanical domains.

2012 Geotechnical Field Investigations

0

5

10

15

20

25

30

35

40

Tonalite VCM VCS Schist Ash Tuff Diorite Felsic Tuff MVC

F F / m

Lithology

Boxplot of FF/m Data by Lithology

Minimum

Average

Maximum




Geotechnical Logging

SRK staff carried out detailed geotechnical logging in order to characterize the rock mass using the In-situ Rock Mass Rating (“IRMR”) method after Laubscher and Jakubec (2000).

Location, orientation and physical properties of large-scale structures; and

Collected/submitted core samples (for site and laboratory strength testing).

During the geotechnical field investigations, quality assurance and quality control (“QA/QC”) was

managed by site visitation of SRK geotechnical staff and several sections were re-logged and compared

to data acquired on site.







Table 16–24: 2012 Geotechnical Drilling Summary

Borehole ID Northing Easting Dip Dip Direction Length Purpose

BH12-SLC-01 751630 196510 -60 115 500

Characterize rockmass betweenshears for pit wall and undergroundstabilityCharacterize cavability of Sericite andTonalite

BH12-SLC-02 751795 196615 -70 15 400

Characterize Rockmass in north wallof pit to help with determiningpermeability and water flow50 m packer intervals

BH12-SLC-03 751740 196870 -70 340 400

Characterize Rockmass in north wallof pit to help with determiningpermeability and water flow50 m packer intervals

BH12-SLC-04 751705 197113 -60 220 550Characterize rockmass around SLCinfrastructure and confirm location ofshears at depth

BH12-SLC-05 751596 197031 -70 0 260Characterize rockmass around SLCinfrastructure

BH12-SLC-06 751406 196711 -60 20 625

Characterize rockmass around SLCinfrastructureconfirm location of shears at depthCharacterize Tonalite and SericiteShears cavability at depth

BH12-SLC-07 751698 196887 -70 215 625

Characterize rockmass around SLCinfrastructureconfirm location of shears at depthCharacterize Tonalite and SericiteSh bilit t d th




deposit. Because these features may affect the stability of the open stope and increased hydrologic

conductivity it decided to include them in the new geomechanical domains (Figure 16-33).

Figure 16–33: Updated geological model with underground geomechanical units.
















Figure 16–38: UCS strengths by Geomechanical Domain

Due to the strong to intense foliation, the laboratory testing results are believed to be influenced by

microdefects and rock fabric (Figure 16-39).

0

50

100

150

200

250

300

Tonalite Mafics Interbedded Sericite QFP

U C

S ( M P a )

Lithology

Boxplot of UCS Data by Lithology

Lower Quartile

Minimum

Average

Maximum

Upper Quartile




Figure 16–40: Distinct strength groupings of the UCS test results.

Table 16–26: Summary for the Material Testing per Underground Domain.




Table 16–27: Geotechnical Data per Underground Domain used for Design Criteria.

Numerical Analysis

Numerical modeling analysis were performed using the 3D finite element continuum code FLAC3D(Itasca, 2009) for stability analysis of the walls created by SLR mining, and mining induced disturbance or

relaxation in the rockmass around the pipe (Figure 16-41). The analysis was carried out by the following

mining sequence:

Saprolite Pit Removal (2 Stages);

Hard Rock Pit Removal (5 Stages);

Sub-Level Retreat Stage Removal (in 50m levels);

Waste Rock Inclusion (if applicable).




pressure distribution created by the mining operations would adversely affect the stability of the SLR

mining excavation and influence future disturbance along the shears (Figure 16-42).




16.3.4 Numerical Modeling Results

Subsidence

Vertical ground disturbance (or subsidence) within the project area was reviewed in order to support

the placement of a dike to protect the mining area from a 1 in 10,000 year river flood events. Results of

the numerical model, as seen in Figure 16-43, suggest that vertical displacement induced by the

underground mining in the north wall of the pipe will be limited to within the weaker sericite schist. The

area beneath the dike is expected to undergo less than 1 cm of vertical displacement which is within the

acceptable limits.




results show influence of mining at maximum distance of approximately 60m from the pipe to the north

and 100m from the pipe in the south. Regions outside of the shear zones were typically only disturbedalong foliation. Colored areas in Figure 16-44 (right) also indicate zones with possible increased hydraulic

conductivity. The modeling results also indicate that the mining induced disturbance would not migrate

along these weaker sericite units.




Figure 16–45: Plan view - Modeled pore water pressure distribution around the mine at an end-life mining

stage at -400m below ground surface.

Mine Infrastructure

Wireframes of the mine design infrastructure were introduced within the numerical model to review the







Table 16–28: Model Sensitivity Results

Ground Support Recommendations

The ground support design criteria and recommendations have been provided for lateral and vertical

development and selected critical infrastructure areas.

Table 16-29 provides a summary of geotechnical parameters used for the ground support assessment.

The majority of infrastructure will be within ground conditions considered to be of good to very good

rock mass quality except development located in the sericite shear which is in category of fair to goodrock quality.

Table 16–29: Underground evaluation rock mass parameters




Table 16–30: Development types and empirical design input parameters




100 mm aperture, #6 gauge weld wire mesh (galvanized) across back and shoulders, (to

within 3.0 m of the floor in Decline, to within 2.75 m of the floor with other capital

infrastructure; and

2.4 m long friction anchors (39 mm, galvanized) installed in walls as required.

Operating Development Headings

For the operating development where spans are between 5.0 m and 6.0 m, the recommended ground

support consists of:

2.4 m long friction anchors (39 mm, galvanized) installed on a 1.5 m x 1.5 m spacing acrossback and shoulders with 150 mm square plates; and

100 mm aperture, #6 gauge weld wire mesh (galvanized) across back and shoulders to

within 3.5 m of the floor.

For development in the saprolite (from portal to fresh rock contact), additional use of 50 mm fibrecrete,

mesh and split set bolts should be applied floor to floor.

For the capital infrastructure, where spans are between 7.0 m and 8.0 m, the recommended groundsupport consists of:

50.0 mm of fibrecrete to within 0.5 m of the floor;

2.4 m long fully grouted rebar (22 mm) installed on a 1.5 m x 1.5 m spacing across back and

shoulders with 150 mm square plates;

7 0 m single Garford cable bolts on a 3m burden and 2m spacing (6m of cable bolt




Cable bolt length should be reviewed on a design span case-basis by the site geotechnical engineer. The

development of intersections in zones of poor ground should be avoided. Breakaways should be

staggered to limit four-way intersections.

Due to the generally good rock quality and shallow depth of the underground mine, detailed stress

analyses are not considered necessary for ground support design at this stage in the project. Stress

analyses can be completed to supplement empirical guidelines provided here to evaluate the potential

extents of stress-induced damage on extraction levels, and anticipate the level of required support.

Ground Control Management Plan

A Ground Control Management Plan (“GCMP”) will be required once underground development

commences at Aurora Gold Project. A GCMP is a live document that is prepared, reviewed and approved

by all key stakeholders It is intended to provide a background on the likely ground conditions, required

procedures, and policy controls in place to manage the risks related to the rock mass conditions. The

GCMP captures key features of the ground control design, implementation, and monitoring. The GCMP

is normally updated annually by the site rock mechanics engineers as conditions change (with

subsequent internal approval), with external reviews completed on a regular basis.

Development of a GCMP is beyond the scope of the Feasibility Study but should be drafted during early

development to establish a common understanding of the ground control standards.

Geotechnical Monitoring Program Recommendations

The monitoring of open pit stability and SLR mine is imperative to follow best practices for both

production and safety reasons A properly established monitoring network provides valuable




Phase 2 – The Final Design would take into consideration the final underground design,

updated geology and structures. The detailed costing and implementation schedules are

included.

Phase 3 - Installation and Commissioning Phase. The provisional design will be reviewed and

updated if material changes occurred in terms of the mine plan. Typically, the

consultant/contractor, together with the instruments manufacturer, would provide

supervision of the drilling program, installation of instruments, and commissioning of the

systems.

Phase 4 – Monitoring and Data Analysis. In this phase, mine personnel are provided with

appropriate training in order to develop the ability to record, analyze and interpretmonitoring results. Typically, an external party would also provide ongoing support and

QA/QC.

Where possible, a design should make provision for the reading of a combination of instruments by

independent and different processes, e. g., an automated logging system, a stand-alone logger, and

manual readings. Instruments and data loggers should be placed with due regard to access in case of

failures and protection from exposure to mobile equipment, blasting, dust and moisture.

Decline Portal Evaluation Program

At this stage, detailed geotechnical drilling has not been completed for the proposed portal location.

SRK would recommend the following studies are completed for the design of a portal box-cut:

Specific geotechnical drill holes characterizing the talus and overburden materials, depth to




The mining methods adopted for the FS are the open benching and SLR methods. Open benching will be

utilized for the first six sublevels and SLR will be utilized for the remaining sublevels. Both open benching

and SLR are top down retreating mining methods applied to steeply dipping deposits with strong host

rocks. The methods do not utilize backfill and are non-caving mining methods. Ore fragmentation is

engineered through blast design and SLR differs from open benching where an ore or low grade blanket

is left over the top of the drawpoints to protect personnel and equipment from fall rock and to minimize

dilution entry.

SLR is similar to sublevel caving (SLC), but without the caved waste behind the drawn ore. The main

differences are in blasting (confined blasting in the case of SLC) and there is a much higher risk ofmudrush events occurring with SLC as more fines are generated and the cave material above the

production levels provides storage for water.

Based on the current underground geotechnical knowledge, SRK considers the open benching and SLR

mining methods appropriate and the most likely methods to deliver an economically viable project with

acceptable operational safety standards and productivities.

16.3.6 Mine Design

The underground mine design was prepared with help of Maptek Vulcan design software and scheduled

using MineRP Enhanced Production Scheduler (EPS). The FS mine design exploits Rory’s Knoll from a

depth of 70 mbsl to a depth of 970 mbsl.

The mine design extracts all Mineral Resource categories (Measured, Indicated and Inferred) above a 1.2

g/t Au cut-off. Only Measured and Indicated categories have been used in the Mineral Reserve estimate.










The mine design is based on rubber tired diesel powered mobile equipment with loader mucking and

truck haulage material haulage. Loaders will dump production material into ore passes where it will be

loaded into trucks by truck loading chutes spaced vertically every 100 m. All life of mine (LOM) capital

infrastructure has been designed outside of the known sericite shear zones.

16.3.7 Cut-Off Grade

The FS design has used a 1.2 g/t Au cut-off grade (COG) for fresh rock mineralized material. The 1.2 g/t

Au cut-off was based on a US$1,300/oz gold price and the preliminary economic parameters used in the

estimate are detailed in Table 16-31 below.

Table 16–31: Underground Cut-Off Grade (COG) Estimate

Items Unit Value

Site Operating Cost

Underground mining US$/t 22.04

Processing US$/t 12.77

G & A US$/t 3.20

Total US$/t 38.01

P i d

Gold price US$/oz 1,300

Gold payable % 99.9




Production Rate Selection

The production rate selected for the orebody was determined by the orebody geometry and continuity,ground conditions, number of drawpoints available, anticipated productivity and scheduling to balance

the resources to achieve a practical production output.

The FS has used a nominal production rate of 1.9 Mtpa and this compares with benchmarked

underground operations utilizing sublevel caving and sublevel retreat mining methods. The production

rate estimated in the FS is within 5% of industry standard empirical guidance criteria (Taylor’s Rule

(Taylor, 1976).

Dilution and Recovery Assumptions

The mine design stope walls have been designed within the limits of the defined mineralization. Stope

wall dilution therefore is confined mainly to the mineralized area and to the northern and southern

bounding sericite shear zones.

Both internal and external dilution sources have been evaluated and accounted for in the dilution

estimate. Material contained within the planned stope shapes below the 1.2 g/t Au cut-off (inferred

material has been assigned a gold grade of 0 g/t) account for the internal dilution sources. Material

outside of the ore development and stope shapes account for the external dilution sources. Low grade

and barren material within the planned stope shapes have been assigned a recovery factor of 40%

where only the swell material is mined and the remaining blasted material is left. This strategy has been

employed to improve the grade factor and to leave the material behind to act as a dilution blanket. Ore

development has been assigned 0% dilution and 100% recovery in order to avoid double counting







Figure 16–51: Stope dilution with increasing in mining depth

Higher recovery estimates have been assigned to open benching sublevels and lower recoveries have

been assigned to lower SLR sublevels in order to maintain an acceptable grade factor and to minimize

dilution entry. Ore recovery factors range from 95% in the open benching sublevels to 90% in the SLR

sublevels (before dilution modifying factors have been applied).

The following summarizes the dilution and recovery modifying factors applied to the schedule to

0%

2%

4%6%

8%

10%

12%

14%

16%

18%

20%

-1000-900-800-700-600-500-400-300-200-1000

Depth (m)

Stope Dilution


h d li ti Th LOM d li d t i 2 4 /d d th l t l



scheduling are conservative. The average LOM decline advance rate is 2.4 m/d and the average lateral

advance rate in all lateral development heading types is 14.4 m/d. A total of 22 km of lateral and vertical

capital development and 35 km of operating lateral development will be completed LOM. The pre-

production and LOM physicals and profiles are summarized in Table 16-32 and Figure 16-52 below.

Table 16–32: Capital and Operating Development Physicals

Type

Development

Profile, Height x

Width (m)

Pre-

Production

Quantity (m)

LOM Quantity

(m)

Capital Lateral Development Meters

Decline,

stockpiles and

sumps

6.0 x 5.5 arch 2,425 10,977

Fresh air access 5.5 x 5.5 arch 173 1,361

Stope access 5.5 x 5.5 arch 498 5,239

Sublevel access 6.0 x 5.5 arch 218 1,639Return air access 5.5 x 5.5 arch 342 1,291

Orepass access 5.5 x 5.5 arch 78 940

Service bay 7.0 x 7.0 arch 0 144

Pump station 7.0 x 7.0 arch 95 475

Capital Vertical Development Meters

F h i 4 0 d 102 1 004




Figure 16–52: LOM capital and operating development physicals

The portal will be collared south of Rory’s Knoll at the process plant ROM pad. The decline will develop

to the north side of Rory’s Knoll where all the sublevel accesses and capital infrastructure will be

developed from. Figure 16-53 below shows the design of a typical sublevel layout.

0

1

2

3

4

5

6

Year

2016

Year

2017

Year

2018

Year

2019

Year

2020

Year

2021

Year

2022

Year

2023

Year

2024

Year

2025

Year

2026

Year

2027

M e t e r s ( 0 0 0 ' s )

Operating Lateral Capital Lateral Capital Vertical




Figure 16–53: Plan view of a typical underground sublevel

The fresh air and egress raise has been designed to be advanced in 100 m vertical raise bore legs and the

exhaust air raise below the 70 mbsl has been designed to be advance in 50 m drop raise legs. The raises

will be advanced as the decline advances. A steel, self-supported enclosed egress ladder way will be

installed in the fresh air raise once all inter sublevel connections have been made.

A sublevel access drive is developed from the decline to within 20 m of the orebody at 25 m vertical

intervals with a sump in the access to collect water from the sublevel and water pumped from the

decline face. A stope access drive is driven perpendicular to the sublevel access drive in both directions

to the limits of the orebody length. Access drives to the orepass and both the fresh air and exhaust

raises are developed followed by completing the orepass, fresh air and exhaust raises.

The remaining operating waste, ore development is completed starting with the middle drawpoint drive.

At the end of each ore drive a slot drive is driven 15 m long perpendicular to the ore drive, defining the

width of each stope. The following figure shows the overall underground development layout plan for

Rory’s Knoll.


16 3 9 Mine Production



16.3.9 Mine Production

The overall stope production sequence is a top down retreating sequence with the top sublevel leadingthe sublevels below it. A minimum 25 m lag between adjacent sublevels stope advance criteria is

maintained for stability and safety. The orebody footprint is large enough to accommodate three

sublevels in concurrent production.

The stope sequence on the sublevel starts with initiating production from the drawpoint in the middle of

the orebody followed by mining adjacent stopes in both directions. A maximum lag of 15 m is

maintained between neighbouring stopes.

Sublevels have been designed to be spaced 25 m vertically and drawpoints are 15 m apart (stope width).

The stope design parameters have been adopted from modern sublevel caving geometries and the

following stope design criteria has been used:

Production stopes

33 m stope height (ore drive backs to the top of the apex pillar);

15 m stope width;

Strike length equal to the orebody width (15 m – 120 m); and

60 degree apex pillar angles.

Production Slots

25 m sublevel spacing;







Figure 16–56: Ore production tonnes and grade inventory

16.3.10 Underground Development and Production Schedule

The LOM lateral development factor (capital and operating) is 450 t of mineralized material per meter of

development advance. The pre-production, development and production LOM schedule physicals for

Rory’s Knoll are summarized in Table 16-33 below.

0

1

2

3

4

0

1

2

3

4

Year

2018

Year

2019

Year

2020

Year

2021

Year

2022

Year

2023

Year

2024

Year

2025

Year

2026

Year

2027

Year

2028

Year

2029

Year

2030

Year

2031

A u ( g / t )

O r e T o n n e s ( 1 0 6 )

Tonnes Grade


Table 16–33: Rory’s Knoll LOM summary of schedule physicals

Underground Schedule Summary Units 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031 TOTAL

Capital Development - Lateral

Decline Capital Development m 1,355 1,070 629 503 1,000 843 784 1,133 799 1,116 717 1,027 10,977

Decline Capital Development kt 121 78 56 45 53 76 70 64 71 63 64 56 818,431

Other Capital Development m 196 1,114 984 1,079 959 871 881 929 888 916 915 738 10,470

Other Capital Development kt 17 96 87 95 84 76 77 82 78 81 80 64 916,213

Capital Infrastructure Development m 0 95 0 0 239 0 30 65 0 95 0 95 619



January 2013 193

Capital Infrastructure Development m 0 95 0 0 239 0 30 65 0 95 0 95 619

Capital Infrastructure Development kt 0 9 0 0 22 0 3 6 0 9 0 9 55,792

Total Capital Development m 1,551 2,279 1,612 1,582 2,198 1,714 1,695 2,127 1,687 2,127 1,633 1,859 22,065

Total Capital Development kt 139 183 143 139 159 152 150 152 149 152 144 129 1,790,437

Capital Development - Vertical

Vertical Capital - Raisebore m 102 204 199 200 197 97 100 199 197 202 203 224 2,124

Vertical Capital - Raisebore kt 4 7 6 6 6 2 3 6 6 6 6 7 65,363

Vertical Capital - Drop Raise m 0 106 157 205 178 177 202 197 212 171 187 127 1,919

Vertical Capital - Drop Raise kt 0 4 6 8 7 7 8 7 8 6 7 5 73,888

Total Vertical Capital Development m 102 310 356 405 375 274 302 396 409 373 389 351 4,043

Total Vertical Capital Development kt 4 11 12 14 13 9 12 13 14 13 13 11 139,251

Operating Development

Development Waste m 652 1,093 1,168 795 723 742 881 756 827 778 649 9,063

Development Waste kt 51 85 91 62 56 58 68 59 64 60 50 703

Development Ore m 1,056 3,096 3,680 2,513 2,018 2,236 2,396 1,983 2,438 2,836 2,112 26,364

Development Ore kt 81 240 285 195 156 173 186 154 189 220 164 2,044

Development Au Grade g/t 2.91 3.52 3.91 3.38 3.49 3.62 3.36 2.48 1.88 2.51 3.06

Development AU Ounces koz 5 22 32 25 17 19 22 17 15 13 13 201

Total Operating Development m 1,708 4,189 4,848 3,308 2,741 2,978 3,277 2,738 3,265 3,613 2,761 35,427

Total Operating Development kt 132 325 376 257 213 231 254 212 253 280 214 2,747

Production Ore

Stope kt 1,271 1,750 1,631 1,722 1,505 1,549 1,459 1,533 1,430 1,519 1,810 2,073 1,955 591 21,797

Slot kt 147 139 143 137 140 144 163 136 125 169 202 195 158 11 2,010

Total Production kt 1,418 1,890 1,774 1,860 1,644 1,693 1,622 1,669 1,555 1,688 2,012 2,267 2,113 603 23,807

Ore Production Au Grade g/t 2.62 2.63 3.21 3.40 3.15 2.98 3.14 3.05 3.36 3.20 2.56 1.83 2.14 2.50 2.82

Ore Production AU Ounces koz 119 160 183 203 167 162 164 164 168 174 165 133 145 49 2,156

Total Ore Mined Summary

Mined Ore kt 81 1,658 2,175 1,969 2,016 1,818 1,878 1,776 1,858 1,775 1,851 2,012 2,267 2,113 603 25,851

Mined Au grade g/t 2.09 2.66 2.75 3.28 3.40 3.18 3.04 3.16 2.99 3.18 3.14 2.56 1.83 2.14 2.50 2.84

Mined AU Ounces koz 5 142 192 208 220 186 184 181 179 181 187 165 133 145 49 2,357


16 3 11 U d d E i Fl



16.3.11 Underground Equipment Fleet

The fleet budgeted and planned to mine Rory’s Knoll at a nominal rate of 1.9 Mtpa is summarized inTable 16-34. Fleet productivities have been based on first principal calculations, benchmarking and

practical experience. Drills, bolters, loaders, trucks, charge machine and support equipment

requirements were based on the estimated required operating hours in each period and the number of

units of each piece of equipment needed to meet those hours.

Table 16

–

34: Underground mobile equipment list

Model Number Type LOMQuantity LOMRebuilds

Max

FleetSize

Sandvik DD421 Jumbo 6 3 3Sandvik DS411 Bolter 6 4 3Sandvik LH517 UG loader 9 8 5CAT AD55B UG truck 28 20 14Sandvik DL421-7 LH drill 2 2 2Cubex Aries ITH drill, V-30 1 1 1Normet Charmec MC605 DA Development charge 2 1 1Normet Charmec LC605 VE Production charge 2 1 1

Normet Utilift MF 540 Services 5 3 2

Veekmas FG 15 C Grader 2 2 1Miller Toyota Hurth Light vehicle 53 39 14Normet Multimec MF 100 Water truck 3 2 1Normet Multimec MF 100 Fuel truck 3 2 1Normet Multimec MF 100 Service truck 3 2 1


The labour estimate assumes a seven-day a week two 12-hour shifts per day operation Technical



The labour estimate assumes a seven day a week, two 12 hour shifts per day operation. Technical

services personnel will work a 2 weeks on, 2 weeks off roster and underground operational personnel

will work a 2 weeks on 1 week off roster. The estimate includes a heavy reliance on expatriate technical

staff, supervision and underground operators. Guyana’s lack of underground mining experience will

require a comprehensive training effort, which is planned to commence during the pre-production

period. As the Guyanese personnel are trained, a reduction of expatriate personnel is planned.


Table 16–36: Technical services personnel

Technical Services 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031

Technical Services Manager 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1



January 2013 196

g

Mine Superintendent 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Maintenance Superintendent 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1Chief Engineer 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Chief Geologist 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Senior Engineer 1 1 2 2 2 2 2 2 2 2 2 2 2 2 2 1

Senior Geologist 1 1 2 2 2 2 2 2 2 2 2 2 2 2 2 1

Mining Engineer 2 4 4 4 4 4 4 4 4 4 4 4 4 4 2

Ventilation Engineer 2 2 2 2 2 2 2 2 2 2 2 2 2 0

Geotechnical Engineer 2 2 2 2 2 2 2 2 2 2 2 2 2 0

Geologist 2 2 6 6 6 6 6 6 6 6 6 6 4 4 4 2

Mining Technician 4 4 4 4 4 4 4 4 4 4 2 2 2 0

Chief Surveyor 1 1 2 2 2 2 2 2 2 2 2 2 2 2 2 1

Surveyor 3 3 6 6 6 6 6 6 6 6 6 6 6 6 6 6

Total 13 15 35 35 35 35 35 35 35 35 35 35 31 31 31 18

Table 16–37: Underground operations personnel

Underground Operations 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031

Supervision 15 15 21 21 21 21 21 21 21 21 21 21 18 18 18 18

Development Operators 36 36 44 47 41 41 41 44 41 41 41 41 0 0 0 0

Production Operators 3 3 65 77 77 80 77 80 83 86 89 92 99 105 102 90

Expatriate Operators 9 9 18 18 18 18 18 18 18 18 18 18 0 0 0 0

Maintenance Personnel 27 27 69 69 72 72 72 72 72 72 72 72 39 39 39 30

Support Personnel 9 9 21 21 21 21 21 21 21 21 21 21 12 12 12 9Contractors 16 16 16 16 16 16 16 16 16 16 16 16 16 6 6 0

Total 115 115 253 268 265 268 265 271 271 274 277 280 184 180 177 147


16.3.13 Material Handling



16.3.13 Material Handling

The mine design is based on rubber-tired diesel powered 17 tonne capacity loaders and 55 tonnecapacity trucks. Trucks will haul material from the underground via the decline to the surface ROM pad.

Ore will be either stockpiled near the crusher or tipped directly into the primary crusher. Waste will be

stockpiled separately and used for site construction purposes. Development waste will be loaded

directly into trucks at underground remuck bays and level accesses. Development and production ore

will be loaded directly into trucks from chutes located vertically every 100 m off of the decline or

sublevel access. Traffic management in the decline will be controlled by block lights and radio

communication with loaded trucks having the right of way.

16.3.14 Mine Safety

Fire Prevention

All diesel equipment (light vehicles and heavy duty mobile equipment) will be equipped with automatic

fire suppression systems and hand held fire extinguishers. Hand held fire extinguishers will be located

throughout the mine at refueling bays, workshops, explosive and detonator magazines, refuge chambers

and lunch rooms. Refueling bays, workshops, explosive and detonator magazines will be equipped with

automatic deluge systems.

A mine-wide stench gas warning system will be installed at the two fresh air intakes to alert

underground workers in the event of an emergency.

Mine Rescue

A mine res e team ill onsist of members sele ted and trained from the A rora proje t site and at


accordance with best practices. Access to the ladder way will be provided on every production sublevel



accordance with best practices. Access to the ladder way will be provided on every production sublevel

and access will be established prior to production mining commencing on the sublevel.

Dust Control

A dedicated water truck will be used for dust suppression of the decline and active sublevels. Spray

nozzles operated by the mobile equipment drivers will be installed at any material loading points for

dust control.

16.3.15 Mine Services

Power

The main project power loads were designed considering the mine equipment, ventilation, air cooling

and dewatering. As the national power grid does extend to the mine site, the power for all of the mining

activities will be generated by power generating plant using a heavy fuel oil diesel gen-sets. A full

description can be found in Appendix E.

Power Distribution

Power for the mine will be supplied by several main feeders sourced from a main 13.8 kV substationlocated on surface. Utilizing 13.8 kV provides the capability to develop much larger development

segments before needing to establish a shorter tie. This allows permanent infrastructure to be

constructed, tested and commissioned well in advance of the next phase of development as shown in

the following Figure 16-57 Single Line Diagram.




January 2013 199

Figure 16 –57 8 kV Surface and Underground Power Distribution – Single Line Diagram


Surface



One feeder will provide power to the main ventilation fans via an overhead power line approximately 1km in length. A small switching substation will be located adjacent to the return air raises. Dual 1.5 MVA

substations (13.8 kV- 4.16 kV) will provide power to the two 1000 kW ventilation fans. Packaged 4.16 kV

variable speed drives will provide the ability to regulate these fans as required to optimize power

consumption.

Two additional feeders also via an overhead power line will terminate at the return air raise switching

substation. These will transition to vertical cable feeders, one located in each return air raise.

Connections into permanent 13.8 kV infrastructure will be made as need.

Underground

The two return air raise feeders will serve as the primary feeders for all of Rory’s Knoll. Ties are provided

at strategic levels to provide a level of redundancy as well as switching capability for maintenance and

construction.

Initial decline and lateral development will be provided with electrical power from a temporary feederuntil the return air raises are developed to a point where cables can be installed in the raises. The

temporary feeder can then be de-commissioned.

The SLR mining method allows the electrical distribution system to be arranged into 9 major portable

substations (1 MVA) each servicing 4 sublevels. Where additional development takes place such as raise

boring and ramp development additional substations can be installed as required. Smaller (500 KVA)


Energy Demand



gy

Energy demand is based on loads identified in the equipment fleet with associated de-rating factors.Factors have been selected based on anticipated equipment duty cycles required to support the mining

method as well as development and production schedules.

Loads associated with dewatering, ventilation and refrigeration have been estimated separately and

included as part of the overall load profile shown in Table 16-38.

Table 16–38: Electrical Energy Demand

Yrs. Ventilation Refrigeration U/G Dewatering Total

Avg. [MW] Avg. [MW] Avg. [MW] Avg. [MW] Peak [MW] Min. MW Avg. [MW] Peak [MW]-2 0.657 0.000 0.198 0.163 0.326 0.989 1.018 1.211-1 0.736 0.000 0.218 0.278 0.555 1.200 1.232 1.5431 1.811 0.000 0.302 0.278 0.555 2.345 2.390 2.7132 2.283 0.000 0.456 0.278 0.555 2.948 3.016 3.3623 2.392 0.962 0.546 0.392 0.785 4.209 4.291 4.7654 2.489 1.131 0.631 0.392 0.785 4.549 4.644 5.1315 2.578 1.301 0.716 0.507 1.014 4.995 5.102 5.717

6 2.658 1.471 0.802 0.507 1.014 5.317 5.437 6.0657 2.731 1.776 0.888 0.622 1.244 5.883 6.016 6.7718 2.784 1.945 0.900 0.622 1.244 6.116 6.251 7.0089 2.833 2.115 0.900 0.737 1.473 6.450 6.585 7.456

10 2.878 2.285 0.900 0.852 1.703 6.779 6.914 7.90111 2.920 2.454 0.900 0.852 1.703 6.991 7.126 8.11212 2.778 2.454 0.900 0.852 1.703 6.849 6.984 7.97113 2.814 2.454 0.900 0.852 1.703 6.885 7.020 8.007


Figure 16–58: Electrical Energy Demand



16.3.16 Ventilation

The purpose of the mine ventilation design is to provide sufficient amount of air for all the mining

activities while satisfying the respective legislative requirements. A full description can be found in

Appendix F.

The ventilation design at Rory’s Knoll for the SLR mining method was carried out for mining down to -

979 Level using VUMA-network simulations program and applying the ventilation criteria indicatedfurther in the text.

It was estimated that about 460 m3/s of fresh air would be sufficient to provide ventilation for all

production and development activities. Fresh air into the active sublevels will be supplied by Fresh Air

Raise (FAR) and distributed to the stopes by auxiliary ventilation system. Exhaust will be carried out by

dedicated sublevels Return Air Raises (RAR). Fresh air to the decline will enter through the decline portal

and after ventilating the decline the exhaust will be carried out by the dedicated decline RAR. Both RARs

will be connected to balance the overall pressure (Figure 16-59).





During the development stage and before the full air circuits are established, all of the development



headings will be ventilated by auxiliary ventilation consisting of auxiliary fans connected to the

corresponding size duct.

Ventilation Criteria

The ventilation criteria for the Aurora underground project were governed by the legislative

requirements, number of operating diesel equipment, daily produced tonnage, mining environment

conditions and sound ventilation practice.

Production rate 5000 t/day

Minimum air volume of 0.06 m3 /s per each kW of diesel power

Air velocity in traveled declines 4 - 6 m/s.

Air velocities in dedicated airways up to 8 - 15 m/s.

Air velocities in dedicated ventilation raises up to 18 - 22 m/s.

Overall mine air pressure drop about 3500 Pa.

Auxiliary fans pressure in flexible duct about 2500 Pa.

Auxiliary fan pressure in rigid duct about 3500 Pa.

Auxiliary ventilation – fan usually one size smaller that the duct.

Ventilation Allocations

The amount of air allocated to a work area is determined mainly on the mining activity and associated

diesel powered equipment. For the Aurora Gold Project ventilation allocations were based on the

b f t k i th d li d th b f j i i i t i th ti bl l T bl


Total Operating Power [kW] 6639



Estimated Airflow Based on 0.06 m /s per kW

Sub-Total Estimated Airflow[m

3/s]

398

Contingency – 15% 60

Total Estimated Mine Airflow

[m3/s]

458

It is expected that only 75 % of the total truck fleet will be engaged in the underground activities at any

one time; the rest of the fleet would travel on surface or undertake the regular maintenance.

Primary Ventilation

Two independent ventilation circuits are proposed for the project. One circuit will be dedicated to the

5.5 m W x 6.0 m H decline and the other one to the production sublevels. Air delivery to the production

circuit would be about 270 m3/s of cooled air a part of which would be supplied as required into the

decline. The decline circuit will circulate about 190 m3/s of uncooled air as the truck fleet and all major

mobile equipment will be equipped with air-conditioned cabins. The airflow for each of the circuits will

be controlled by the exhaust fans positioned at the top of the return airways.

The decline circuit intakes air through the decline portal and exhaust it via the decline Return Air Raise

(RAR). Breakthrough into the exhaust raise will be every on every level about every 100 meters of the

vertical interval. Supply air will always flow freely to the lowest RAR breakthrough and from there

carried to the decline advancing face by the auxiliary system until the next breakthrough to the RAR is

completed. Some cool air will enter the ramp on -54 level via a crosscut connecting FAR to the ramp.


Long-lateral development to access the sublevels; and



Short lateral development to access and ventilate the active faces.

The decline development will be ventilated by twin flexible duct in the forced fresh air supply mode in

1000 m segments. The exhaust will free flow back to the decline portal until the first breakthrough into

the decline RAR is reached (end of the first 1000 m twin duct segment at about 150 m vertical interval).

From there the air will be exhausted via the RAR. The duct will now move to develop the second

segment of the decline. The whole process will be repeated until the whole decline is fully developed.

The forced air mode was selected as it provides more effective ventilation to the face and there is no

need for scavenger vent system to boost the face airflow velocity.

The sublevel drives development will be ventilated either as the ramp extension or a branch off, until

the system of ventilation raises is completed with the sublevels breakthroughs.

Fresh air for the development leading to the active faces will be ducted from the sublevel drives and

exhausted free flow into the exhaust raises.

Pump stations, sub-stations and other service ventilation centres will be ventilated in series with the

fresh air delivered to the production zones. Maintenance workshops will be ventilated in parallel to thefresh air and returned directly to the RAR.

Development Fans & Duct Specifications

The purpose of the auxiliary ventilation for development or active headings is to take air from the main

air stream and distribute it to the particular workings.


at each of the decline RAR cross-cuts. As the decline progresses the pressure in the decline circuit will



increase.

The sublevel circuit will receive fresh and cooled air from the FAR and after traversing the sublevel the

air will exhaust through the RAR. Connection to the FAR and RAR will be provided at every sublevel.

Ventilation regulators will be provided at the RAR connection on each sublevel. Typically these would be

roll-up garage type doors, adjustable louvers or guillotine type regulators operated by chain blocks. The

flow will be regulated in range of 0 m3/s to 40 m3/s. System of airlocks will have to be constructed to

separate the decline and sublevel ventilation circuits.

In both circuits the supply air volume will gradually increase until the full production stage has been

reached.

Main Fans – Specifications

The size of the main fans was calculated on the proposed openings sizes and volume of the supply air.

Two 2800 mm diameter axial flow fans each for one of the proposed ventilation circuits with duties of

230 m3

/s @ 3400 Pa equipped with 1000 kW motors will control the project airflow.

Each fan will be equipped with inlet box turning vanes and evase. A floating shaft will be used to connect

the drive to the impeller. Two-speed motors, adjustable pitch in motion or Variable Frequency Drive

could be used to control the required air volume through the development and production phases of

the operation. See Figure 16-60 for the main fans General Arrangement.




January 2013 208

Figure 16 –60 Exhaust Fans - General Arrangement


16.3.17 Air Cooling



g

Cooling required was estimated from heat load balances taking auto-compression, rock heat, vehicleheat and all other components into account. On those bases the size of the refrigeration plant was

determined as 11 MWR and the air cooling will be delivered by two 5.5 MWR (Megawatt of

Refrigeration) conventional refrigeration plants. For a detailed description, refer to Appendix F. There is

a possibility to utilize the waste heat from the power generating sets in the absorption chillers. One unit

of the absorption chiller utilizing the available waste heat can produce about 1.2 MWR and it could be

used to augment the conventional plant either in midterm of mining to defer the purchase and

construction of the second conventional unit or to increase the refrigeration capacity

The mine refrigeration system will be installed on surface and the complex will comprise a BAC,

refrigeration modules, plant building, cooling towers, water and pump systems and electrical and

control systems. The general arrangement drawing for the refrigeration system is illustrated in Figure

16-61.




January 2013 210

Figure 16 –61 Air Cooling plant – General Arrangement


The refrigeration and air cooling facility will operate automatically and will be monitored remotely

i h h d f i i i h b d i h i f



without the need for permanent on-site operators. Provision has been made in the cost estimate for

appropriate monitoring and control systems.

Air Cooling Design Criteria

The following criteria were used to determine the size of the refrigeration plant.

Surface Ambient Design Condition

Summer design wet-bulb [wb]/dry-bulb temperatures [oC] 26.0/31.4

Barometric pressure [kPa] 101.0 Surface air density [kg/m3] 1.14

The above values are based on hourly averages of measurements made at the project site over the

periods 2006 to 2009 and 2011.

Geothermal Properties

Applied thermal properties and the geothermal gradient.

Virgin Rock Temperature [VRT] can be described by:

VRT [°C] = 23.77 + [0.01169 x depth below collar (m)]

The VRT between -79 and -979 levels ranges from 24.7 oC to 35.2 oC.

The recent thermal properties are indicated in Table 16-40.


Ramp Conditions



The maximum wet-bulb temperature in the ramp will be limited to 32 °C provided the trucks in the rampare fitted with enclosed air-conditioned cabins and may not be exceeded except in emergency

conditions. If a truck breaks down it would be repaired/salvaged by personnel who would have to work

for short periods in a high temperature atmosphere. The maximum dry- bulb temperature will be 37 °C.

In the event of vehicle failures/breakdowns personnel should only be allowed to work outside the air-

conditioned cabs for short periods in these elevated temperatures.

Estimated Heat Loads

A full interactive computer simulation using VUMA3D-network software was applied to determine air

temperatures, flow rates, heat loads and cooling requirements. The simulation considers the following

heat load components.

Surrounding rock – exposed rock in all openings

Broken rock – blasted rock

Auto-compression – conversion of potential energy into enthalpy, increases with depth

Diesel powered equipment – conversion of combusted fuel to heat

Auxiliary fans – conversion of electrical energy into heat energy

Ground water – when exposed to air and at the temperatures higher than VRT

Other sources – such as pumps, electrical sub-stations, workshops, lighting, explosives,

strata movement, etc., assigned values from experience based on the size of operation and

production rates.




Water Pumping System



For the proposed mining layout the pumping infrastructure required for normal operation the pumpingsystem will have the following pumping sections:

Main pumping system capable of pumping the required volume of water from the mine.

Development pumping system: to collect water from the decline development and pump to

the main pumping system.

Local pumping to collect water from the development face to pump to the main pumping

system

The arrangement of the pumping system will be dependent upon the location of the lowest main pump

station in relation to the decline development. It is possible that additional submersible pumps will be

required to assist with pumping to the lowest main pump chamber should the water produced in the

decline be excessive.

The pump selection criteria were based on required flow rate, operating pressure and water quality.

Single stage pumps rather than multiple stage pumps are proposed. Although they have a lower delivery

head they could handle dirty water and are more suitable for this type of mining operation. To minimize

the number of pump stations only pumps with high head capacities are considered. In addition to this a

number of pumps are connected in series in each pump set. The proposed dewatering method and

layout is shown in Figure 16-62.




January 2013 215

Figure 16 –62 Mine Dewatering System Schematic


Pumping Volumes



Pumping volumes have been calculated based on the average flow requirements resulting in averageflow of 27 l/s. In flood conditions the required flow over a 24 hour period is 357 l/s. This is more than

thirteen times the average flow.

Pump Selection

The pump selected for this application is the Weir DWU dewatering pump. The pump would have a

nominal flow rate of 100 l/s (360 m3/h) and a delivery head of 80 m to 85 m. The flow duty is

conservative and makes allowances for a loss in pump efficiency and potential changes to the pumping

layout.

The proposed configuration requires a pump station to be constructed every six levels (150m vertical

lift). Three pumps are to be connected in series for each pump set. A 200NB pipeline would be installed

in the decline. One pipeline is required for each pump set. The pipeline would deliver to the next dam

up the decline or to the surface if pumped from the upper pump station.

The development pumps will be vertical spindle pumps installed inside a tank mounted on a skid. This

unit can be moved as the decline development progresses. The development pump unit transfers waterdelivered from the face pump at the lowest portion of the decline to the main pump or the next

development pump station in the decline. The development pipeline would use 100 mm HDPE pipe.

Normal Pumping Operations

During normal pumping operations the average required flow rate is 27 l/s. The main pumps will only be


Steel piping with flanged connections is used for the main pump pipelines as this type of piping is best

suited for permanent installations for high resistance to potential damage



suited for permanent installations for high resistance to potential damage.

The main pumping system uses 200NB piping to match the flow requirement of 100 l/s per pump set.

The operating pressures are 2,345 kPa for steady state and 3,000 kPa for the surge pressure.

16.3.19 Communications

The communications and tracking systems will consist of the following:

Communications Backbone

The communications backbone in the decline and connected to the surface control room will be a single

mode multi pair fiber optic cable. The protocol will be TC/IP Ethernet. A fiber optic cable will also link all

PLC’s in the motor drive MCC’s with network switching devices (OTN and Ethernet switches) to link the

networks.

Leaky Feeder Network

The leaky feeder system will allow communications on the levels and will also be used as the transport

system for the man and vehicle tracking system. Communication will be by two ways radios issued topersonnel and mounted on vehicles.

Data and voice communication throughout the mine will be achieved through the use of an Ethernet

supported “leaky feeder network”. This consists of antenna-cabling being installed throughout the main

drives in the mine, with signal repeaters and boosters installed at periodic locations.


Personal Emergency Devices (PEDs)



All personnel will be issued with PEDs. A separate radio communication system will be installed for thissystem.

Operator Control

A central control room located on surface is proposed to control all the major aspects of the

underground operation, such as loading stations, ore passes, ventilation system, service bay and

communication.

16.3.20 Explosives Storage and Initiation SystemThe primary explosives storage area will be located on surface. Secondary facilities will be located

underground to supply up to 7 days of explosive usage.

Two explosives storage facilities are planned underground on the -170 and -545 levels to separately

store bulk explosives and detonators. Each storage facility will be located off of the decline in a

dedicated return air way. They will be equipped with fire suppression, concrete floors, waste water

sumps and secure steel locked gates.

Centralized blasting will be utilized in the underground mine to initiate development and production

blasts. The proposed system is the BlastPED firing system (i.e. Mine Site Technologies). Explosive

handling, loading and initiation are to be carried out by trained and authorized personnel using the

owner’s procedures and industry best practices.

16 3 21 Fuel Storage and Distribution


16.3.23 Water Supply



Service Water System

Service water for the underground operation is used mainly for drilling, dust control, workshops,

washing and fire suppression of class “A” fires. The water will be supplied from the service water tank

located at the vicinity of the decline portal and gravity fed into the decline pipeline. From there it will be

distributed to the sublevels via installed water pipes.

Pressure reduction valves will be installed as needed. Spray nozzles operated by the mobile equipment

drivers will be installed at any material loading points for dust control.

Potable Water

Potable Water will not be supplied to the underground mine by a separate piping system. Instead,

potable water will be delivered to each refuge station, lunchroom and also carried by mine personnel.

16.3.24 Equipment Maintenance and Service Bay

An underground work shop with two service bays will be constructed to provide for routine

maintenance of the loaders and drills. The service bay(s) will be constructed at the -340 sublevel. It isenvisaged that the service bay(s) will be used for routine service, and minor maintenance and repair

work on scoops and drills. Beside the service bay(s), a wash-down bay would be provided equipped with

an oil trap. The oil will be collected in the dirty oil tank while the water will report to the drainage

system. One service bay will be equipped with an overhead crane plus general tools and equipment.

Ventilation of the service bay welding area will be connected directly to the mine exhaust system.


Table 16–41: Aurora Combined Open Pit and Underground Production Schedule

Aurora Combined Production Summary Units 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031 Total

Open Pit Production Schedule

SAP Ore Mined dmkt 28 858 - 154 1,804 1,475 463 179 1 160 5,123

Fresh Ore Mined dmkt - 894 1,755 1,842 1 346 876 1,186 1,176 477 8,550

Total Open Pit Ore Mined dmkt 28 1,752 1,756 1,996 1,805 1,821 1,542 1,484 1,682 637 13,673

Open Pit Mined Grade Au (g/t) 1.06 2.33 2.90 2.67 1.59 1.90 3. 24 2.17 2.93 4.19 2.55

Open Pit Mined Ounces koz 1 131.3 164.0 169.6 92.3 111.5 147.5 96.0 125.6 85.8 1,120



January 2013 220

Sap Waste Mined dmkt 69 2,507 90 105 4,388 8,179 6,455 3,101 65 1,460 26,421

Fresh Waste Mined dmkt 3 1,364 4,040 1,973 81 1,997 6,690 9,854 9,071 2,243 37,318

Total Waste Mined dmkt 72 3,871 4,130 2,078 4,469 10,176 13,145 12,955 9,136 3,703 63,738

Strip Ratio (W:O) 2.6 2.2 2.4 1.0 2.5 5.6 8.5 8.7 5.4 5.8 4.7

Waste Rehandle dmkt 18 72 72 72 73 73 73 73 73 73 676

Stockpile Reclaim dmkt - - 1 - - - 20 3 119 505 - 828

Underground Production Schedule

Total Underground Ore Mined dmkt - 81 1,65 8 2,175 1,969 2,016 1,818 1,878 1,776 1,858 1,775 1,851 2,012 2,267 2,113 603 25,851

Underground Mined Grade Au (g/t) - 2.09 2.66 2.75 3.28 3.40 3.18 3.04 3.16 2.99 3.18 3.14 2.56 1.83 2.14 2.50 2.84

Underground Mined Ounces koz - 5 142 192 208 220 186 184 181 179 181 187 165 133 145 49 2,357

Underground Waste Mined dmkt 142 193 155 153 172 161 162 165 164 165 158 140 - - - - 1,930

Combined Open Pit and Underground Production

Total Ore Mined dmkt 28 1,752 1,756 2,077 3,463 3,996 3,511 3,500 3,500 2,515 1,776 1,858 1,775 1,851 2,012 2,267 2,113 603 39,524

Total Mined Grade Au (g/t) 1.06 2.33 2.90 2.65 2.10 2.36 3. 26 2.88 3.06 3.33 3.16 2.99 3.18 3.14 2.56 1.83 2.14 2.50 2.74

Total Mined Ounces koz 1 131 164 175 234 304 355 316 312 269 181 179 181 187 165 133 145 49 3,477


1 7 . 0 R E C O V E R Y M E T H O D S



0 C O O S

PR O C E S S S T R A T E G Y 17.1

The proposed process encompasses two separate phases to accommodate the processing of three

separate ore types; Saprolite, open pit Fresh Rock, and underground Fresh Rock. Phase 1 consists of

crushing followed by a single ball mill grinding circuit providing an operating capacity of 5,000 k/d. Phase

2 incorporates a second ball mill installed in parallel to the Phase 1 ball mill and increases operating

capacity to 10,000 k/d. The process downstream of the milling circuit will be similar in both phases.

The Aurora gold processing facility is designed to treat a nominal 1.75 Mt/a during Phase I and 3.5 Mt/a

after Phase 2 construction is complete. Both the Saprolite and the Fresh Rock ore types are readily

amenable to conventional cyanide leaching. The milling circuit will grind and leach 1.75 to 3.50 Mt/a of

ore respectively to produce gold doré. The milling plant will have overhead cranes and a roof. The gold

leaching process will use a modified CIL circuit for leaching and recovery to carbon. All of the processing

uses industry proven processes.

In both phases Fresh Rock will be crushed prior to the single stage ball mill grinding section followed by

thickening, leaching, CIL, carbon desorption, and eluate electrowinning. The Saprolite will be processed

in conjunction with Fresh Rock whenever available. Upon the depletion of the Saprolite ore the process

facility will then begin treating 100% fresh rock.

G ld d é ill b d d i h i fi d d i l i i


Table 17–1: Process Design Criteria

Criteria Units Design



Criteria Units Design

Crusher Availability % 70Crusher Throughput Phase 2 t/h 595.2

Primary Crusher SelectionType C145

No 1

Secondary Crusher SelectionType HP400

No 1

Tertiary Crusher SelectionType MP800

No 2

Mill Throughput Phase 2 Mt/a 3.5

Mill Availability % 95Mill Throughput Phase 2 t/h 452.9

Physical Characteristics (fresh rock)

BWI kWh/t 14.4

RWI kWh/t 16.0

CWI kWh/t 16.2

SMC (A x b) kWh/m 35.8

Specific Gravity t/m 2.80

Grind Size µm 109Head Grade (design) g/t Au 3.33

Gold RecoverySaprolite % 94

Rock % 90Preleach Thickener Rate (fresh rock) t/m /h 0.6Preleach Thickener Underflow Density (fresh rock) % w/w 45Leach Circuit Type - CILLeach Circuit Residence Time h 24Leach Stages Phase 2 - 2Carbon Adsorption Stages Phase 2 - 6Detoxification Process - Air/SO2


Ease of operation and maintenance



The selection of these parameters is discussed in the following sections.

17.3.1 Throughput and Availability

The process comminution circuit includes a three stage crushing circuit followed by one 18 ft x 26 ft ball

mill for each phase. The grinding mill is sized to achieve the design throughput of the fresh rock ore

types, with the potential for increased throughput whenever saprolite ore sources are available. The

overall process facility availability of 92% is an industry standard for similarly sized milling circuits with

moderately abrasive ore. The major comminution design parameters are the following:

Crusher Work Index (CWI) of 16.2 kWh/t based on the 75th percentile of the samples tested

at SGS;

Bond Rod mill Work Index (RWI) of 16.0 kWh/t based on the 75th percentile of samples

tested at SGS;

Bond Ball mill Work Index (BWI) of 14.4 kWh/t based on the 75th percentile of the samples

tested at SGS;

Bond Abrasion Index (A

I

) of 0.346 g based on the 75th percentile of the samples tested atSGS;

Target grind size P80 of 109 μm based on various leach tests programs.

Sizing of the crushing and grinding circuits was determined through evaluation of comminution Test

work performed by SGS Mineral Services (SGS). The ore characterization test work provided data on

Crusher Work Index Bond Ball and Rod Mill indices Bond Abrasion Index and SMC and JK A x b values




January 2013 224

Figure 17 –1 Tertiary Crushing and Reclaim PFD


PR O C E S S D E S C R I P T I O N 17.5



The process facility design is based on a flowsheet with unit process operations that are proven in theminerals processing industries. Material handling of the Saprolite ore can be difficult due to the in-situ

moisture, fine particle size, and cohesiveness if not handled properly. To mitigate the risk of downtime

when handling this material well known industry engineering practices were incorporated in the design

and sizing of equipment. The Aurora Project gold circuit includes the following unit processes:

Dump pocket for Run-of-Mine ore (RoM). RoM from the open pit and underground mines

will be crushed using a primary jaw crusher to a product size of nominally 80% passing (P80)

115 mm. The crushed ore will be fed onto a conveyor that contains a self-cleaning magnetto remove excess steel, and the ore will be transferred to the secondary crusher feed

conveyor;

A single secondary cone crusher of 315 kW operating in open circuit;

A single scalping screen ahead of the secondary cone crusher;

Two tertiary cone crushers (one per phase) of 600 kW with operating in closed circuit;

Two double deck inclined vibrating screen to close the flow of ore around the tertiary cone

crushers;

The crushed ore will be collected in a conical stockpile with a live capacity of 16 hours. Ore

from the stockpile will be reclaimed using two apron feeders, each capable of feeding 125%

of the full mill throughput;

Two 4.3 MW ball mill with dimensions of 18ft x 26ft, operating in closed circuit with 375mm

hydrocyclones;

Preleach thickening of cyclone overflow in a 30 m diameter high compression thickener to


Raw water will be reclaimed by a barge from the raw water dam. The raw water is recovered and

distributed throughout the site.

Potable water will be generated on-site by the treatment of a fresh water source using multimedia filter,



chlorine, ultra filtration and ultraviolet light. Potable water will then be distributed for use in the processfacility and around the site.

Other items included in the process include Process instrumentation, control devices, and process

facility and instrument air services including associated infrastructure.

PR O C E S S D I S C U S S I O N 17.6

17.6.1 Reclaim, Saprolite Handling, Crushing, and Ball Circuit

Three stages of crushing will prepare the feed for grinding. The crushed ore will feed onto a covered

crushed ore stockpile.

Crushed ore will be reclaimed from the ore stockpile using two apron feeders. The apron feeders will

discharge the ore onto the Ball mill feed conveyor. Each apron feeder has been designed to deliver 125%

of the designed mill feed rate.

Ball mill feed weightometers will be located on the mill feed conveyor(s). The weightometers willprovide feed rate data for control of ore reclaim feeders.

Ore is fed into the mill feed chute and discharged from the ball mill trommel (rotating screen) before

reporting to the cyclone feed pump box. Oversize pebbles or steel from the rotating screen (scats) will

be collected and recycled back into the mill feed conveyor or discarded to the TMA.


Water recovered from the preleach thickener will be pumped back to the grinding circuit for density

control.

Leaching of gold by cyanide will occur in a hybrid CIL circuit, comprising two leach tanks and six CIL



tanks. The Phase 1 leach circuit will operate with one leach tank and three CIL tanks. The CIL circuit willprovide a total leach residence time of 24 hours. Each CIL tank will have two intertank pump screens to

enable slurry flow while maintaining the fluid level near the top of the steel tank. Two intertank screens

are supplied to ensure proper carbon (charcoal) management and satisfactory flow rates during times of

variable slurry viscosity associated with the saprolite feed. An overhead gantry crane will enable the

removal of intertank screens for maintenance and cleaning. Carbon pumps will transfer carbon between

CIL tanks and then transfer loaded carbon to the elution circuit.

The sodium cyanide solution will be automatically dosed to the leaching circuit and controlled by an on-line cyanide analyzer via operator inputs. Cyanide levels will be maintained during the initial stages of

leaching, and allowed to decay during the latter stages, minimizing cyanide in the tailings. The on-line

cyanide analyzer will measure both feed and tailings cyanide levels to minimize excess cyanide being

used. Manual samples will be collected by the leach operator to supplement the online analyzer.

Slaked lime will be added to the grinding and leaching circuits to maintain pH levels. Compressed air will

be sparged into the CIL tanks to maintain required dissolved oxygen (DO) levels. Hydrogen cyanide

(HCN) gas detectors will be located on top of the leach tanks and tailings areas. An additional HCNdetector will be located in the cyanide detoxification area.

17.6.4 Carbon Desorption and Regeneration

Loaded carbon will be pumped from the CIL circuit to the loaded carbon screen, located above the acid

wash column. Slurry will be washed from the carbon, which will be fed by gravity to the acid wash


17.6.5 Refinery

Pregnant solution from the desorption circuit will be pumped through three parallel electro-winning

cells, located on a mezzanine floor in the refinery.



The electro-winning cells will recover gold onto stainless steel cathodes. Once sufficient gold has beenloaded, the electro-winning cells will be taken off-line and cleaned. A high-pressure water spray cleaner

will be used to wash the plated gold off the cathodes. The gold containing sludge will be filtered and

placed into a retort for drying and mercury removal.

Silica, borax, nitre and soda ash will be combined to form a flux, which will be used to remove impurities

during smelting. The fluxes will be weighed out according to the desired mix, and combined using a flux

mixer. The flux will be combined with the dried gold sludge and smelted in the diesel-fired furnace.

The doré will be weighed on a set of precision scales and placed in a vault to await transportation off-

site. Slag produced during the smelting operation will be collected in a bin, and recycled into the ball

mill.

17.6.6 Tailings Detoxification

Leached slurry from the CIL circuit will gravity flow over the carbon safety screen and will be pumped to

the cyanide detoxification (detox) circuit. The detox circuit is designed to reduce weak acid dissociable

cyanide (CNWAD) to less than .5 ppm. The circuit will consist of two agitated detox tanks with a total

residence time of 2.5 hours.

A 20% (w/w) sodium meta-bisulphite (SMBS) solution (a source of SO2) will be added to the

detoxification feed slurry. Oxygen in the form of compressed air will be sparged into the detox tank to

maintain a high redox potential to maximize oxidation of the cyanide present. Slaked lime will be dosed


Process water;

Potable water;

Fire water.



Raw water will be collected from the Fresh Water Pond via a raw water barge to service the processfacility, and the general buildings. The raw water pontoon pumps will deliver raw water 1 km to the raw

water tank located on the process facility site. Raw water will be used for process make-up

requirements, gland seal water, dust suppression, reagent mixing and fire services.

Process water is defined as water that is internally recycled within the process facility site footprint. It is

comprised of decant (reclaim) water from the TMA plus pre leach thickener overflow water.

Potable water will be sourced from the Fresh Water Pond and rain water harvesting. The water will bepumped to a water treatment facility located near the process facility site. The water will be treated and

stored in the potable water tank. Potable water will be distributed throughout the process facility.

Potable water will be used for all eye wash stations and safety showers.

Firewater, a subset of raw water, will be connected throughout the process facility, laboratory,

workshop and fuel storage areas via dedicated firewater pump system, which includes a back-up diesel

pump. The raw water tank will have a dedicated reserve for firewater.

17.6.9 Reagents

Logistics and security of supply will be an important management function throughout the mine life due

to the project location. Individual commodities are discussed below.

Sodium Cyanide


The slaked lime will be stored in an agitated storage tank, where it will be recirculated around the

process facility in a ring main system using the lime ring main pumps (duty/standby arrangement). The

lime addition to grinding, leach and detoxification will be controlled automatically using pH probes and

control valves. Redundant pH monitoring will be used in the leaching and detox area.



Activated Carbon

Activated carbon will be delivered to the site in 500 kg bulk bags. Activated carbon will be stored in the

reagents storage area of the warehouse facility and transferred as required in the CIL circuit.

Activated carbon will be loaded by hoist into the carbon quench tank located in the carbon regeneration

area. The carbon will be washed to remove any fines before being pumped to the CIL circuit. Fines will

be washed from the new carbon to minimize potential gold losses to the tailings.

Sodium Metabisulphite

SMBS will be supplied in 1000 kg bulk bags as a dry reagent. SMBS will be used as a source of SO2 for the

cyanide destruction circuit (air/SO2 process).

SMBS bulk bags will be lifted by an overhead hoist and loaded into the mixing tank by way of a bag

splitter. Dilution water will be added to produce a solution concentration of 20% (w/w). The diluted

solution will be transferred to the SMBS storage tank. Metering pumps (duty/standby configuration) will

dose SMBS to the NaCN detox circuit as needed.

Copper Sulphate

Copper sulphate will be supplied in 1000 kg bulk bags as a dry reagent. Copper sulphate is a chemical

used as a catalyst for the cyanide destruction circuit (air/SO2 process). Copper sulphate will be stored

on site in the reagents area of the warehouse facility Copper sulphate bulk bags will be lifted by an


circuit. Spent acid may be recycled four times (or until the level of contaminants is considered too high)

before being disposed of in the CIL tailings pump box.

Flocculent



One flocculent mixing, storage and dosing system will be provided adjacent to the preleach thickeners.Dry powder flocculent will be mixed with raw water to make 0.25% w/v solution in a packaged

flocculent mixing system. The mixed flocculent solutions will be pumped to a storage tank with 8 hours

capacity at design flow rates. Flocculent solution will be dosed to each thickener by duty/standby,

variable –speed helical rotor pumps. Process water will be mixed into the flocculent lines to further

dilute the flocculent solution to 0.05% w/v before it will be added to the thickener feed slurry.

Fluxes

Sodium borate, more commonly known as borax, along with silica flour, soda ash and potassium nitrate

will be delivered to the site on a pallet containing 25 kg bags, and unloaded using a forklift. The fluxes

will be stored in the warehouse, and transported to the gold room as required.

Grinding Media

Forged carbon steel grinding media will be delivered to the site in 20 t containers. It is anticipated at

start up 75 mm to 88 mm balls will be used for charging.

Ball mill media will be delivered into a bulk ball storage bunker. Media addition will be controlled by

adjustments made by the process facility control room operator, who will modify the set point in the

process control system (PCS).

An overhead crane in the primary milling area will be used to load steel balls into the ball mill. Steel balls

will be transported from the ball mill ball storage bunker by a front end loader to a small ball loading


outputs to alarms, control the function of the process equipment, and provide logging and

trending facilities to assist in analysis of process facility operations; and.

Uninterrupted power supplies will provide operating control stations with 20 minutes of

standby power.



The general control strategy adopted for the Aurora Project gold process facility is as follows:

Integrated control via process control system (PCS) for areas where equipment requires

sequencing and process interlocking;

Hardwired interlocks for safety of personnel;

Motor controls for starting and stopping of drives at local control stations, via the PCS or

hardwired depending on the drive classification;

All drives can be stopped from the local control station at all times. Local and remotestarting is dependent on the drive class and the control mode;

Control loops via the PCS except where exceptional circumstances apply;

Monitoring of all relevant operating conditions on the PCS and recording select information

for data logging or trending; and

Trip and alarm inputs to the PCS will be fail-safe in operation (i.e., the signal reverts to the

deenergized state when a fault occurs).

Drives that form part of a vendor package will be controlled from the vendor’s control panel. At a

minimum, “Run” and “Fault” signals from each vendor control panel will be made available to the

SCADA system via the PLC. Where practical, the PCS will interface with the vendor control panel to

provide full operating status, including state of all drives, alarms, and instrument outputs.

PR O C E S S F A C I L I T Y I N F R A S T R U C T U R E A N D S E R V I C E S17 8


PR O C E S S F A C I L I T Y A N C I L L A R Y B U I L D I N G S 17.9

Ancillary buildings will be required for the operation and maintenance of the process facility. These

include the laboratory, chemical and reagent storage, and process facility control room. The warehouse

will be housed in the adjacent vehicle maintenance and administration building. There will be



will be housed in the adjacent vehicle maintenance and administration building. There will be

mechanical, electrical, instrumentation and general items in designated areas. Internal offices will be

supplied for warehouse and maintenance staff.

The process facility control room will be located within the grinding circuit of the process facility. A

sampling laboratory located within the process facility site will provide the requirements of the process

operations. Additional equipment will be provided to allow basic metallurgical investigations.

E Q U I P M E N T S I Z I N G 17.1017.10.1 Primary Crushing

RoM will be dumped from haul trucks, or a front-end loader (FEL), through an 800 mm square-grid static

grizzly into the 100 t capacity RoM hopper. Ore will be recovered from the RoM bin by an apron feeder

feeding a vibrating grizzly, to remove fines from crusher feed. The grizzly oversize material reports to the

1.4m x 1.1m single toggle jaw crusher.

17.10.2 Secondary and Tertiary CrushingThe primary crushing discharge conveyor will deliver the primary crushed ore to the secondary crushers.

Screen oversize material will discharge to standard cone crushers (secondary crusher) with closed side

settings of approximately 19 mm.

The combined secondary crusher discharge product and screen undersize would be transported by a


The crushed ore stockpile will have a live capacity of 6650 t. The total capacity of the stockpile is

approximately 16 hours to one and one-half days of nominal ball mill feed capacity depending on Phase

1 or Phase 2 of the operation.

The stockpile ahead of the ball mill will be covered with a light weight cover to enhance the material



The stockpile ahead of the ball mill will be covered with a light weight cover to enhance the material

handling issues.

17.10.4 Grinding Circuit

The grinding circuit is designed to process 10,000 t per day of RoM. To account for variations in ore

competency, the 75th percentile of measured comminution parameters derived from samples tested

were used for the circuit design. This would indicate that the current circuit could achieve higher

throughputs when processing less competent ore.

The grinding circuit includes a three stage crushing circuit followed by two 18 ft x 26 ft diameter ball

mill. The grinding mill is sized to achieve the design throughput for the fresh rock ore types; with the

potential for increased throughputs should Saprolite ore sources be available during Phase 2.

A power-based approach was used for the grinding mill sizing. This approach uses empirically derived

models developed from a database of actual process facility operations data and associated bench-scale

testwork. Critical input parameters to the model are ore competency (measured by either JK drop

weight A x b or SMC DWI values) and Bond Work indices (crushing, rod and ball). The power basedmodel predicts the milling efficiency of the various circuits based on JK drop weight/SMC data.

The installed ball mill power of 4,300 kW (per mill) incorporates allowances for drive train losses as well

as a design contingency to account for the accuracy of the models, calculations and test work used to

determine the expected average pinion power.


17.10.7 Cyanide Detoxification

Two 1208 m3 detox tanks were selected to provide 2.5 hours of retention as indicated during laboratory

testing.




1 8 . 0 P R O J E C T I N F R A S T R U C T U R E



PR O J E C T L O G I S T I C S 18.1During construction and mine operations, a combination of transportation methods, including aircraft,

river navigation, and road access will be needed to supply the project. Overall, the offsite infrastructure

will include docking facilities for cargo ships at the Buckhall Port facility on the west side of the

Essequibo River. The mine access road will be 150 km in length from Buckhall Port to the mine site.

During construction, the Aurora Gold Project will require substantial efforts to mobilize equipment,

materials, and workers from Georgetown to the site and vice versa. Some existing off-site infrastructurewill require improvements to allow efficient and safe development of the mine. These improvements

need to commence at the end of each wet season to ensure successful completion.

During production, the main road will be mainly used for the supply of food, reagents, spare parts,

mining supplies, and diesel fuel. The site airstrip will be used mainly for personnel transportation and

emergency situations.

ON -S I T E I N F R A S T R U C T U R E 18.2The site entails a series of open pits, waste rock stockpiles, a process facility with associated laboratory

and maintenance facilities; maintenance buildings for underground and open pit equipment. Facilities

and structures include a warehouse, office, change house facilities, ventilation shaft, mine air cooling

process facility, explosives storage area, power generating station, fuel storage tanks, a warehouse and

laydown area a 1 200 m airstrip and a permanent accommodation complex The open pit area will be


The power plant’s generating sets will generate power at 13.8 kV, 60Hz. The process facility’s main

electrical room will be fed with two 13.8 kV lines from the main power plant in order to ensure full

redundancy. All other loads of the project will be fed at 13.8kV from the power plant through overhead

power lines. These power lines will also be used to deliver power to various locations to support



activities during the construction of the project.

The power plant will have its own fuel storage facility. Given the remoteness of the Aurora Gold Project

site and its accessibility, a one month on-site fuel storage capacity for No 4 Fuel Oil will be provided to

accommodate for continuous operation of the power station. Two 1,000 m3 fuel oil storage tanks will be

built next to the power plant.

18.2.2 On Site Roads

Project site roads include haul roads suitable for use by mining trucks and service roads for use bysmaller vehicles. The site roads will be built for use only by authorized mine personnel and equipment,

with access controlled by Guyana Goldfields.

Saprolite covers the entire project area. Roads will, as far as practical, be constructed using cut and fill

techniques to achieve design alignment and grade. Placed saprolite fill will require compaction in small

lifts in order to provide a competent road foundation. At several locations the compacted saprolite fill

will serve as a dike to divert surface water drainage and protect the mining areas from water ingress.

Both haul roads and service roads will require a surface layer of crushed rock fill to facilitate all-season

use. The rock fill will generally be sourced from the project rock quarries, since some roads are

scheduled to be constructed during the pre-production period when minimal fresh waste rock is

available from the open pits. Road surfacing material will break down in time, and frequent re-surfacing

during the mining operation will be required as part of an ongoing road maintenance program. Dust


This haul road is designed for the single lane 40 t articulated truck traffic, with pullouts to permit truck

passing.

Approximately a 15 to 20 km network of the service roads will be built to provide the access to

environmental discharge points, airstrip, explosive storage facility TMA, MWP, and the underground



g p , p, p g y , , g

shaft, locations.

18.2.3 Utilities and Services

Fresh Water Supply, Fire Suppression Water and Distribution

Raw water will be sourced from water wells drilled into the underlying bedrock, collection of surface

water from creeks and rain water harvesting systems. A rain water harvesting area will be constructed at

the mancamp to supply fresh potable water and fire suppression water for building services such as

dining facilities, showers and toilets. An in-line chlorine metering system will disinfect the water supply.

Potable water for the process facility and operations and maintenance facility will be obtained from a

roof collection system on the operations and maintenance building.

Fresh raw water supply will be obtained from a fresh water pond about 1 km south of the process

facility. This water is primarily for fire protection, make-up requirements for the process facility,

fluidization and flushing for the gravity concentrators, cooling the drives and lube systems, use in the

strip solution heat exchanger, reagent preparation, and gland water distribution.

Sewage Collection and Disposal

A sewage treatment process facility will be constructed just east and downhill of the mancamp site.

Buried sewer pipes will collect sewage from the site to the treatment process facility. The treatment

process facility will consist of two independent containerized treatment lagoon systems working

d d l d d d d f b h d f h d


Communications and IT Systems

Point-to-point satellite communication will be the main communication system between the mine and

the outside world. The system includes voice/data/video/fax, internet, and VPN services, including bi-

directional links between the mine site and Georgetown.



A backup/emergency satellite system will be provided for redundancy. The backup/emergency system

includes voice/data/fax, TV and internet access for a minimum number of users.

VHF/UHF radio communication will be available within a 10 km radius from the process facility. The

phone system will be a voice over internet protocol. This will reduce wiring costs and allow voice-

messaging integration with e-mail. End-to-end IP video connectivity with business quality transmission

will provide video conferencing capabilities. At least three satellite phones installed at strategic areas

will be provided for emergency communications.

Satellite TV for entertainment, cellular communication, and FM radio will be provided.

A cellular phone system from Buckhall Port to the site will be installed. This system will be a joint effort

of Guyana Goldfields, Barama Logging Company and the government of Guyana.

The IT system will be based at the operations and maintenance building and connected throughout the

site by a fiber optic network. The connection between IT devices and end-users will provide high-

throughput, secure, reliable and redundant service for data and voice. The network system will be

connected to protocol independent multicasts (PIMS) and business networks through routers with

firewalls and will provide remote access as required. The system will have security and encryption to

prevent unauthorized access.

Vehicle Fueling Facility and Mine Equipment Ready Line


The mine operations area, approximately 1,200 m2 in size, will house the following: mine operation staff

office, maintenance staff office, heavy equipment/high-rack storage warehouse on the ground floor,

low-rack storage warehouse on the mezzanine floor, first aid room, lunch room, locker room and toilets.

The equipment maintenance shop is designed to repair and maintain the mine fleet and other mobile



q p p g p

equipment. It will consist of four bays for heavy mobile equipment repairs and maintenance, two bays

with two lifting hoists dedicated for heavy vehicle maintenance, two bays allocated for a machine shop,

tire servicing, and other major repairs. A light vehicle maintenance building will be located adjacent to

the operations and maintenance building.

A 50 t bridge crane will be provided for two bays and storage area. A separate truck wash station,

equipped with a washing system with a water/oil separator for heavy mining equipment, will be

installed outdoors.

Mancamp

The permanent accommodation complex will be constructed on a 10 ha elevated site southeast of the

mine complex. The accommodation complex will incorporate the following dormitory styles:

Type A dormitories will be private, single-occupancy rooms;

Type B facilities will be semi-private and have single-occupancy rooms with two rooms

sharing one shower and toilet room; Type C dormitories will be double-occupancy rooms with a central shower and toilet facility

shared by 30 rooms.

The accommodation complex will also include the following facilities:

Kitchen dining hall;


Airstrip

The airstrip will be upgraded prior to dike construction to provide for personnel access, transportation

of sensitive equipment, and medical emergencies. The new facility will consist of a runway and storage

for emergency and firefighting materials.



Incoming and outgoing flights will be scheduled for daylight hours only. Temporary lighting will be

employed along the airstrip in the event of night time medical emergencies. The runway will be 1,200 m

long x 30 m wide with 90 m runway end safety areas at each end. The elevation of the airstrip and

related access road will be above the flooding level for continuous serviceability during flood seasons.

Aircraft maintenance and fuelling will be performed in Georgetown; thus, no provision for aviation fuel

storage facilities will be provided.

Dust suppressants will be used on the runway as required to reduce dust emissions during periods of

little or no precipitation.

Solid Waste Disposal and Recycling Facility

Non-recyclable, non-toxic solid waste will be disposed of in an onsite lined landfill. Used tires will be

shredded and placed in the landfill.

18.2.5 Tailings Management Area

The ore from the process facility will be processed on-site and disposed of as non-segregating slurry in

the TMA. The TMA is located about 1 km southwest of the process facility and is designed to contain

mine life tailings production of 25 Mt.

Tailings will be delivered through a 2 km long pipeline in Phase I to the TMA. Tailings will be delivered


An emergency spillway will be located at the southeast corner of the facility. The spillway inverts will be

raised as the dam is raised through the mine life. The final dam crest will be at 75.0 m elevation with low

flow spillway at 73.0 m. The inflow design flood is the probable maximum flood (PMF). The high flow

spillway will provide a 0.5 m freeboard during the maximum PMF water level. A freeboard of 0.5 m will



be available throughout each stage of dam raise over the maximum PMF water level.

18.2.6 Fresh Water Pond

The fresh water pond (FWP) will be located immediately east of the TMA and will hold approximately

600,000 m3 at the NHWL. The water will be obtained from a 140 ha drainage basin. Overflow from the

FWP will discharge through a spillway into the mine water pond. The water will be used for make-up

requirements for the process facility, fluidization and flushing for the gravity concentrators, cooling the

drives and lube systems, use in the strip solution heat exchanger, reagent preparation, and gland water

distribution. The water will be pumped from a barge mounted pump through a 1 km pipeline to the

process facility.

18.2.7 Mine Water Pond

The mine water pond (MWP) will be located immediately south of the open pit mine and will have a

capacity of 750,000 m3 at NHWL. The MWP will receive water from a 113 ha drainage basin, the 110 ha

open pit mine drainage basin and groundwater inflow to the open pit mine. The MWP is designed to

detain water for a minimum of 7 days prior to discharge. Discharges form the MWP will flow through aconcrete box culvert spillway to a tributary of the Cuyuni River.

18.2.8 Emergency Discharge Pond

The emergency discharge pond will be located immediately west of the process facility. This pond will be

a double lined facility used on an emergency basis if a vessel in the process facility needs to be emptied.


OF F -S I T E I N F R A S T R U C T U R E 18.3

18.3.1 Buckhall Port

Buckhall is the logistics hub for the Aurora Gold Project and is located on the west bank of the Essequibo



River about 24 km up river from the Atlantic Ocean. The facility will include a wharf, pier, bulk fuelstorage, barge slip, customs clearing area, and laydown/staging areas. The facility will include a

government-run customs entry port. This site also includes an administration building, two story

dormitory buildings, kitchen/dining hall, vehicle maintenance and security. Some equipment will be

assigned permanently to this facility. There is a light vehicle maintenance facility. A heavy vehicle

maintenance shop will be constructed for the contract hauling fleet. The site has been fenced and

topped with security razor wire to deter unauthorized access.

The pier will accommodate up to 3,000 t sea-going cargo vessels and landing for barge vessels that willtransship from sea-going vessels too large to travel up the Essequibo River. The fuel depot includes two

existing 94,000 L steel diesel fuel storage tanks, one 1-million L tank for diesel and one 2-million L tank

for No. 4 fuel. Each fuel type will be transferred from incoming fuel barges to the storage area through a

series of dedicated pipes and flexible hoses. Both areas include spill containment berms constructed of

concrete and masonry with an underlying HDPE liner. The containment volume of the spill containment

berm is set no less than 1.5 times the volume of the tanks.

Water is provided from a well and supplemented with rainwater collected from the roofs. Sewage istreated in septic tanks and discharged to leach fields in accordance with Guyanese design regulations.

Diesel generators produce electrical power. The communication system includes locally available mobile

phone service and satellite dish internet service.

Additional facilities to be added at the site include:


constructed and is currently being maintained by Barama Company Limited for logging operations.

Barama will continue to use these sections of the road during its ongoing logging activities.

An improvement plan has been developed that includes a land survey of the road centerline to establish

road stationing. The improvement plan will be drafted for review and coordination with Barama and



Guyana Goldfields. The road will require a wearing surface. Drainage structures and log bridges will be

replaced and upgraded within the first 5 years of operations.


1 9 . 0 M A R K E T S T U D I E S A N D C O N T R A C T S



M A R K E T S19.1Gold metal markets are mature, with many reputable refiners, and brokers located throughout the

world. The advantage of gold, like other precious metals, is that virtually all production can be sold in

the market. As such, market studies, and entry strategies are not required.

Metallurgical process studies confirm that the Project will produce doré of a specification comparable

with existing operating mines.

Demand is presently high with prices showing remarkable increases during recent times. The 36-month

average London PM gold price fix through December 2012 is $1,485/oz.

C O N T R A C T S 19.2

Currently there are no material contracts in place other than those disclosed in this document. However,

Guyana Goldfields has obtained quotes for future service needs. It is anticipated that the following

contracts will be in place upon Project commencement:

Secure doré transportation to market;

Doré refining;

Supplier and service contracts including;

o Barge transportation of supplies to Buckhall Port;

o Diesel and fuel oil;


2 0 . 0 E N V I R O N M E N T A L S T U D I E S ,P E R M I T T I N G A N D S O C I A L O R



C O M M U N I T Y I M P A C T

E N V I R O N M E N T A L A N D S O C I A L S T U D I E S A N D I M P A C T S 20.1

Guyana Goldfields has committed to the establishment of environmental and social practices for the

Aurora Gold Project that comply with the legal requirements established by the nation of Guyana, as

well as current World Bank/International Finance Corporation (IFC) Performance Standards1 (IFC, 2012)

and the IFC “Environmental, Health, and Safety Guidance for Mining”2 (IFC, 2007). Compliance with

applicable IFC Performance Standards is required for all lending actions undertaken by the IFC, as well as

for all projects funded by Equator Principles Financial Institutions (EPFIs) in those nations (including

Guyana) which are not currently designated as Organisation for Economic Co-operation and

Development (OECD) countries. The EPFIs currently include over 70 major private banks with

international operations.3

In keeping with the requirements of IFC Performance Standard 1, “Assessment and Management of

Environmental and Social Risks and Impacts” (PS-1), Guyana Goldfields has undertaken a wide range of

environmental (or combined environmental and social) studies in recent years to assess the social and

environmental impacts likely to be associated with the Project. These studies have included an initial

Environmental and Social Impact Assessment (ESIA) conducted to IFC standards in 2010 (referred to

hereafter as the “ERM ESIA”)4, as well as a separate ESIA conducted for Guyana Goldfields in accordance

i h G E i l P i A (EPA) i b G d S E i i


IFC Public Health Technical Assistance Program for Guyana Goldfields, Phase 1 (Newfields,

2008);

IFC Public Health Technical Assistance Program for Guyana Goldfields, Phase 2 (Newfields,

2009); and

f h l d l



Bio-assessment of the Cuyuni River near Aurora: Environmental and Economic Implications(Dr. Nicole Duplaix, October, 2009).

These studies document field work undertaken since 2006 by several teams of national and

international biologists, ecologists, and social scientists. Taken collectively, these studies confirm that

the environment associated with the Aurora Gold Project ’s area of influence has been significantly

impacted by artisanal and small-scale mining (ASM), logging, hunting, and other human activities for

well over a century. The Cuyuni River has likely served as a transportation corridor since the prehistoric

arrival of the first indigenous peoples in the region. The immediate area of the Aurora Gold Project sitewas first explored in the 1930s, and has been impacted by ASM activities, both legal and illegal, ever

since. Apart from supporting a major logging concession (operated by Barama Company Limited), the

construction of the Barama Road has contributed to a significant increase in human activities in the

region to the north of the Cuyuni River and to the west of the Essequibo River. The Buckhall Port and the

access road corridor to the Aurora Gold Project site have also been subject to frequent disturbances

associated with human activities.

It is noteworthy that the large species of fauna that are otherwise common in pristine habitats alongsimilar types of rivers in this part of South America have been observed to be absent or very rare in the

Project AOI. The absence of such species is a key indicator of historical human impact, presumably due

to the pressures of hunting and the increased turbidity and other degradation of river quality from many

years of logging and ASM activities, as well as from the continuing disturbances created by motorized

equipment and sporadic motorboat and roadway traffic.


to the east of the site. The Cuyuni River originates in Venezuela and extends some 750 km east to the

Essequibo River in Guyana, covering an area of approximately 53,500 km2 (AMEC, 2009).

Surface water quality in the Cuyuni River and its tributaries has, in general, been impacted historically by

upstream ASM activities, particularly in Venezuela, and ASM-associated mercury contamination of



surface water, sediments, and fish has been documented as a concern throughout the Guianas. ASM

activities have also notably increased surface water turbidity and concentrations of suspended solids in

the Project AOI. In 2006 and 2007, surface water samples were collected from locations on the Cuyuni

River, Gold River, and from an unnamed tributary of the Cuyuni River, both upstream and downstream

of the Aurora Gold Project area. Surface water samples from the unnamed tributary may be considered

indicative of background water quality for the many smaller creeks that crisscross the Aurora Gold

Project site. During the 2006-2007 sampling activity, it was observed that total iron was the only

parameter that exceeded the guidelines defined in the IFC “Environmental Health and Safety StandardGuidelines for Mining” (IFC, 2007).5 In 2009, additional surface water and sediment sampling was

conducted along the Cuyuni River to assess baseline surface water and sediment quality conditions.

Total suspended solids, iron, and oil and grease were detected in surface water samples at

concentrations above (IFC, 2007) guidelines. The detection of oil and grease at elevated levels in one of

the samples may be attributed to residual petroleum hydrocarbon impacts from previous gold

exploration and drilling operations near the Aurora Gold Project site. In 2011, surface water samples

were also recovered from five locations around the site. Sediment samples were also recovered from

stream beds at points coincident with surface water sampling locations. Both surface water andsediment samples were tested, and none of the parameters sampled exceeded (IFC, 2007) standards.

Groundwater flow and quality; based on groundwater monitoring conducted by AMEC in 2006 and

2007, shallow groundwater exists within the unconsolidated overburden from approximately 1 to 4 m

below grade. In 2011, a well installation program was initiated to fill the data gaps required to better


Soils; the native soils of the proposed mine site, along the access road, and at the Buckhall Port

generally consist of residual material derived from weathered acidic crystalline rocks (i.e., granite, schist,

dolerite, granodiorite, and phyllite), and alluvial sediments derived from stratified and unconsolidated

deposits of sand, silts, and clays. In upland areas, soils consist of deep, well-drained, yellow and reddish-

b d l l d ll l I i i d ll i l f il l d b



brown sandy clay loams and gravelly clays. In riverine and alluvial fan areas, soils are also deep but rangefrom poorly- to excessively well-drained clay loams and sands (Braun and Derting, 1964).

Geochemistry ; a testing program was conducted for Guyana Goldfields in 2011 and 2012 by Klohn

Crippen Berger (KCB), in order to determine the potential acid rock drainage (ARD)/mineral leaching

potential of representative samples of predicted Aurora Gold Project overburden material. Results of

static and kinetic humidity cell testing indicate that the majority of the samples tested have very low

ARD/metals leaching potential.6 These results will be used to develop a geochemical block model that

identifies areas that are potentially acid generating (PAG), acid generating (AG), and non-acid generating(NAG). This model and associated testing data have played a key role in the development of the

Project’s water quality management strategy, which is discussed in Section 20.3.4. Geochemical data will

also be used adjust estimates of leachate loads to support water quality modeling and the ongoing

management of the Mine Water Pond (MWP), Fresh Water Pond (FWP), Tailings Management Area

(TMA), and TMA Diversion Ponds 1 and 2, with respect to maintaining effluent/seepage water quality

within IFC and International Cyanide Management Code (ICMC) guidelines.

Flora; the Aurora Gold Project concession, the Buckhall Port, and access road corridors were allcompletely forested prior to initiation of regional ASM and logging activities in the region. This part of

Guyana does not support natural open savannah areas or marshes. Canopy trees are the dominant plant

strata, followed by lower-story trees, plants, and undergrowth. None of the plant species identified are

locally endemic, of significance to local communities, or listed as threatened by the International Union

for Conservation of Nature and Natural Resources (IUCN) “Red List” 7or equivalent national or regional




Project

ComponentPotential Impact Management/Mitigation Strategies

Mine and mill/

process facility area

Loss of aquatic

habitats

Installation of diversion structures to route un-impacted surface water

around mining operations, and to route all impacted water to the Mine

W t P d (MWP) F h W t P d (FWP) d T ili M t



Water Pond (MWP), Fresh Water Pond (FWP), and Tailings ManagementArea (TMA)

Implementation of BMPs in the Erosion Protection and Control Plan and

Water Management Plan to manage topsoil/overburden stockpiles; and to

detect and mitigate erosion in other disturbed areas

Implementation of appropriate progressive restoration and erosional

stabilization procedures for mined-out areas of the open pits, as well as

areas disturbed by ASM, per the Erosion Prevention and Control Plan and

Mine Reclamation and Closure Plan

Implementation of a routine water quality monitoring program in theCuyuni River and its tributaries as described in the Water Management

Plan and ESHS Monitoring Plan

Loss of terrestrial

habitat and flora

Minimization of clearance actions/project footprint per the Erosion

Protection and Control Plan, Exploration Management Plan, Early Works

Construction Management Plan, and Construction Management Plan

Implementation of a routine biodiversity monitoring program per the

Biodiversity Management Plan and ESHS Monitoring Plan

Implementation of specific mitigation measures for the protection of any

identified sensitive species and habitats, per the Biodiversity Management

Plan


Water Management Plan to manage sediment generation from waste

rock/topsoil stockpiles; and to detect and mitigate erosional conditions in

other disturbed areas.

I l t ti f i t i t ti d i l


Project


Slope failure of

waste rock and

Saprolite stockpiles

Operator training programs/ compliance with Overburden Management

Plan

Periodic monitoring of the physical integrity of the waste rock and Saprolite



Saprolite stockpiles,disrupting surface

flows

Periodic monitoring of the physical integrity of the waste rock and Saprolitestockpiles, in accordance with the Overburden Management Plan and the

ESHS Monitoring Plan, and regrading and/or strengthening of earthworks or

other action as indicated by observed conditions

Modification of

hydrologic flow

patterns within

streams/creeks due

to FWP, MWP, TMA

and surface/underground mining

operations

Installation of diversionary structures/diversion of un-impacted surface

water around mining and processing operations in order to maintain

biological base flows in Cuyuni River tributaries

Control of discharges from the FWP, MWP, and TMA into tributaries of the

Cuyuni in accordance with the Water Management Plan and Tailings Area

Management Plan, in compliance with effluent discharge guidelinesand/water quality standards defined by Guyana EPA, (IFC, 2007), and the

ICMC

Mine and mill/


Breaches and

overtopping of the

FWP, MWP, and/or

TMA

Provision of sufficient freeboard in the design of the MWP, FWP, and TMA

based on the Probable Maximum Precipitation (PMP) event

Inclusion of sufficient contingency in the design of the MWP, FWP, and TMA

embankments to withstand PMP events plus an appropriate safety factor

Inclusion of a series of redundant water management features (e.g.,

spillways, diversion ponds) in the TMA design

Rigorous independent Construction Quality Assurance (CQA) oversight of

MWP, FWP, and TMA embankment construction

Development and implementation of probabilistic water

balance/monitoring program and other BMPs for the MWP, FWP, and TMA,

in accordance with the Tailings Management Plan, Water Management

Pl d C id M Pl


Project


Mine and mill/


Potential runoff or

seepage of

contaminated water

Installation of barge and pumpback systems to return TMA reclaim water

back to the process facility for industrial use

Installation of embankment seepage interception collection and return



contaminated waterfrom the TMA into

surface water

Installation of embankment seepage interception, collection, and returnsystems on TMA, as described in Section 20.3.

Construction of TMA in a low-permeability saprolite soil basin, supported

by geological evaluation of the basin and local compaction, grouting, or

other basin preparation actions during construction as necessary to ensure

consistent low-permeability conditions, as described in Section 20.3.

Inclusion of a series of redundant water management features (e.g.,

spillways, diversion ponds) in the TMA design.

Implementation of probabilistic water balance/water monitoring program

for the TMA in the operational phase, in accordance with the TailingsManagement Plan, Water Management Plan, and Cyanide Management

Plan, as discussed in Section 20.3.

Regular monitoring of TMA water quality in accordance with the Tailings

Facility Management Plan, Cyanide Management Plan, and ESMS

Monitoring Plan to ensure that controlled discharges will be within Guyana

EPA and (IFC, 2007) limits, as well as the free cyanide limits recommended

by the ICMC for protection of aquatic life

Diesel oil spill into

the Cuyuni River

Implementation of the secondary containment and engineered spill

prevention and control measures, remote fuelling control procedures, oily

water separators/treatment systems, and other BMPs per the Hazardous

Material Management Plan and Spill Prevention, Control and Contingency

Plan

Implementation of the Project’s preventive maintenance (PM) and field

inspection programs for the operation of the fuel farm, emergency

t d f lli t ti


Project


Potential runoff or

seepage of leachate

from the Solid

Collection and periodic testing of leachate from landfill; if testing results

indicate effluent quality issues with respect to Guyana EPA or (ICMC, 2007)

guidelines route to MWP for dilution and storage or install local water



from the SolidWaste Landfill guidelines, route to MWP for dilution and storage or install local watertreatment system

Infiltration of

potential spills or

discharges of

cyanide and other

chemicals into

groundwater

Purchase of cyanide exclusively in solid briquette form, transported in

dedicated stainless steel ISO delivery/mixing tanks

Implementation of the secondary containment, engineered spill prevention

and control measures, and other BMPs defined by the Project Cyanide

Management Plan and Emergency Preparedness and Response Plan

Implementation of operational practices in the process facility that

minimize the potential for process upsets, as noted in the Project Cyanide

Management Plan For other (non-cyanide) reagents and fuel, implementation of the

secondary containment and engineered spill prevention and control

measures, remote fuelling control procedures, oily water

separators/treatment systems, and other BMPs defined by the Hazardous

Material Management Plan, Buckhall Spill Contingency Plan, the Aurora site

Spill Prevention, Control and Contingency Plan, and the Emergency

Preparedness and Response Plan

Mine and mill/


Potential infiltration

of surface waterfrom the Cuyuni

River into open pit/

underground mine

Construction of the Cuyuni River dike system, including seepage collection

wells will be installed

Potential failure of

the TMA and MWP

dams after

Stabilization, breaching/removal of embankments, closure, and selective re-

vegetation of tailings surfaces and embankments per the final Detailed

Mi R l ti d Cl Pl


Project


Buckhall-Aurora

Access Road

Loss/ degradation of

aquatic habitats


Water Management Plan to detect and mitigate areas of soil erosion,

manage stormwater runoff and control sedimentation on access road



manage stormwater runoff, and control sedimentation on access roadROWs and other adjacent disturbed areas

Implementation of progressive restoration and erosional stabilization

procedures for any excessively wide ROW areas, as well as any ASM-

disturbed areas, per the Mine Reclamation and Closure Plan

Implementation of routine water quality monitoring program at stream

crossings per the ESHS Monitoring Plan

Loss/alteration of

terrestrial habitats

Minimization of clearance actions/project footprint per BMPs in the

Exploration Management Plan, Early Works Construction Management

Plan, and Construction Management Plan Implementation of a routine biodiversity monitoring program within the

Aurora concession per the Biodiversity Management Plan and ESHS

Monitoring Plan

Implementation of progressive restoration and stabilization procedures for

ROWs and ASM-disturbed areas, per the Exploration Management Plan,

Early Works Construction Management Plan, Construction Management

Plan, and Mine Reclamation and Closure Plan

Implementation of specific mitigation measures for the protection of

sensitive species and habitats, as directed by the Biodiversity Management

Plan


Water Management Plan to manage stockpiles, and to detect and mitigate

erosional issues in other disturbed areas

Buckhall-Aurora Impacts to Soils Implementation of the Project Erosion Prevention and Control Plan,


Project


Potential runoff or

seepage of leachate

from the Soliddf ll

Collection and testing of leachate from landfill; if test result so indicate,

installation of treatment system prior to controlled discharge



from the SolidWaste Landfill

Buckhall Port Spills of fuel and

other chemicals

being loaded/

unloaded at the Port

Implementation of the remote fuelling control procedures and other BMPs

per the Hazardous Material Management Plan and Buckhall Spill

Contingency Plan

Spill kits for all heavy vehicles operating within the Buckhall compound

Use of portable floating booms and spill response kits around the Buckhall

fuelling terminal and implementation of the secondary containment and

other BMPs at the fuel storage facility, per the Hazardous Material

Management Plan and Buckhall Spill Contingency Plan Accidental

discharges of fuels,

oils and grease from

equipment and/or

from the failure of

fuel containment

facilities

Use of spill response kits and implementation of the other BMPs applicable

to the fuel storage facility, per the Hazardous Material Management Plan

and Buckhall Spill Contingency Plan

Overall macro-scale

impacts of theproject

National socio-

economic impactsdue to closure

Implement the Project Community Relations Plan, Influx Management Plan,

and final Detailed Mine Reclamation and Closure Plan; key actions willinclude:

provision of appropriate retrenchment compensation for the mine

workforce, as described in the and final Detailed Mine Reclamation and

Closure Plan

diversification of skills/training, and building capacities of former

k d i id t fi d i t iti ith


Project


Potential ASM issues Continued consultation and engagement with illegal and artisanal miners in

accordance with the Influx Management Plan and Community Relations

Management Plan



Management Plan.

Potential human

influx to the mine

site and related risks

Train the security forces on site to monitor potential influx, and to handle

influx without creating conflict or security issues, in accordance with the

Influx Management Plan and applicable sections of the Community

Relations Management Plan

Partner with the Government and/or Barama to plan for the sustainable

development/ growth of the Buckhall community, per the Influx

Management Plan and applicable sections of the Community Relations

Management Plan Consultation with the Government on promotion of planned regional

development

Prohibition of onsite hiring. Work opportunities advertised and controlled

through GGI’s Georgetown office

Control of employee travel to and from site using GGI or GGI contractor

vehicles

Prohibit public access to the Solid Waste Landfills

Promote and implement health awareness and disease prevention

campaigns, especially for malaria suppression and humanimmunodeficiency virus/acquired immunodeficiency syndrome (HIV/AIDS)

and sexually transmitted diseases (STDs) among workers and contractors,

as well as local communities upstream and downstream of the project site

Overall macro-scale

i t f th

Potential influx into

B kh ll d t P t

Train the security forces on site to monitor potential influx, and to handle

i fl i h i fli i i i d i h h


S I T E W I D E W A T E R B A L A N C E 20.2

An initial water balance for the site was developed using the GoldSim™ software platform. A model was

prepared to simulate reservoir and pond operational levels from predicted precipitation and other

project inflows and outflows. A conceptual representation of the water balance is shown in Figure 20-3.

Initial water balance results indicate that excess water will need to be discharged to the environment



Initial water balance results indicate that excess water will need to be discharged to the environmentfrom the TMA and the MWP on a continuing basis. Water discharged from the TMA will first pass

through the (dilution) reservoir at Diversion Dam 2 prior to entering the environment. Water from the

MWP will be further diluted by discharge from the FWP prior to discharge to the environment.




Figure 20–1: Conceptual Model, Site Water Balance


MWP for settling and dilution prior to controlled discharge to the environment; see Sections 20.2 and

20.3.4. The stockpile sedimentation ponds and/or MWP will also be used to retain haul road runoff. All

sedimentation ponds will be designed to provide the required one day retention time for runoff from a 2

year, 24 hour storm.

Additional details on waste rock/overburden and Saprolite stockpile design are presented in Section



Additional details on waste rock/overburden and Saprolite stockpile design are presented in Section

16.0.


Runoff and tailings supernatant water will have a minimum retention time of 30 days prior to being

discharged, under mean annual precipitation conditions. Over the first four years of operation, when all

tailings production is deposited within the TMA, the mixing ratio for tailings water and precipitation

within the TMA capture area is estimated to be 1:9 during an average year.

Water quality in the TMA MWP and FWP will be monitored via a surface and groundwater monitoring



Water quality in the TMA, MWP, and FWP will be monitored via a surface and groundwater monitoring

well-based sampling program to ensure that any seepage and controlled discharges will be within the

effluent discharge guidelines and water quality standards defined by the Guyana EPA and (IFC 2007).

Indications of any negative water quality trends may prompt the installation or activation of additional

clarification ponds, treatment and/or polishing systems, or biotoxicity testing studies to demonstrate

that aquatic species are adequately protected.

Monitoring of groundwater and surface water leaving the project site will also be conducted using a

network of monitoring wells and surface water monitoring stations. Upgradient monitoring wells and at

least one upgradient surface water monitoring station will be established to collect background water

quality data.

20.3.3 Site Monitoring

Site monitoring requirements will be managed within the context of a documented ESMS for the Aurora

Gold Project that is designed to comply with the requirements of IFC PS-1, as previously noted. The

continuous monitoring of environmental performance is a key component of ESMS design, and formsthe basis for managing or mitigating the environmental, occupational health and safety, and social

impacts identified by the ESIA process, over the entire project life cycle. Many of the management plans

included in Guyana Goldfields’ ESMS are designed specifically for this purpose, and will contain or

reference specific environmental (and environmentally-linked social) monitoring practices and SOPs.

Major components of the environmental monitoring program collectively represented in these


biotoxicity testing studies to demonstrate that aquatic species in any receiving waterways

are adequately protected.

Erosion Prevention and Control Plan; the Erosion Prevention and Control Plan will be

applied to all areas that are disturbed by mining operations and that are susceptible to

erosion. This will include weekly monitoring of diversionary structures, access and haul road

ROWs the Project airstrip and other recently reclaimed or disturbed areas for erosional



ROWs, the Project airstrip, and other recently reclaimed or disturbed areas for erosional

conditions or turbid runoff that could impact local streams or tributaries of the Cuyuni River,

and the ongoing implementation of various appropriate mitigation measures or BMPs as the

situation may require.

Water Management Plan; as noted in Section 20.3.4 below, Guyana Goldfields will develop

and maintain a probabilistic water balance and supporting monitoring program for the TMA,

MWP, FWP, and associated water management structures. Pond water levels, associated

flowrates, bathymetric data, and site precipitation and other weather conditions will bemonitored and input to the water balance model in order to ensure the maintenance of

adequate freeboard in all seasonal conditions. Water quality in the TMA, MWP, and FWP

will also be monitored to ensure that any seepage and controlled discharges will be within

the effluent discharge guidelines and/water quality standards defined by the Guyana EPA

and (IFC, 2007). Indications of any negative water quality trends may prompt the installation

or activation of additional clarification ponds, treatment and/or polishing systems, or other

studies to demonstrate that aquatic species in any receiving waterways are adequately

protected. Cyanide Management Plan; Guyana Goldfields will develop and maintain a comprehensive

Cyanide Management Plan based on the ICMC standards of practice9 that addresses all

aspects of cyanide procurement, transportation, and management. In the operational phase

of the project, the Cyanide Management Plan will reference applicable operations and

maintenance manuals and other operational SOPs and inspection and preventive


management of accumulated precipitation, segregation of incompatible and materials, the

condition of warning or access control signage, and the functionality or readiness of PPE and

fire suppression and other emergency systems.

Spill Prevention, Control, and Contingency Plan and Buckhall Spill Contingency Plan; both

of these management plans will contain requirements for weekly inspections of

containment integrity management of accumulated precipitation and the functionality or



containment integrity, management of accumulated precipitation, and the functionality or

readiness of PPE, fire suppression systems or equipment, and other emergency systems.

Air Quality Management Plan; the Air Quality Management Plan will require visual

monitoring of roadway dust generation (at the Aurora and Buckhall sites and the Tapir

Crossing) in dry seasons, as the basis for a dust suppression program using water and/or

appropriate non-toxic surfactants. The Air Quality Management Plan will also establish

minimum requirements for an initial ambient workplace air quality survey in the vicinity of

the pit and process facility, as well as the underground mine in later phases of mineoperation. Results of this survey will be used to establish a respiratory protection baseline

for work areas containing potential respiratory hazards.

Noise & Vibration Management Plan; Guyana Goldfields will monitor the noise and

vibration associated with routine blasting operations and provide data back to the Project

Blasting Engineer for use in improving the efficiency of blasting operations. Ambient

workplace noise will be monitored near drilling and loading operations, heavy vehicles, the

mill/process facility, power plant, and other noise-generating areas of operation, as the

basis for designation of hearing protection zones for workers. Additional workplace testingwill be conducted as the basis for establishing hearing protection requirements for specific

job assignments (e.g., chainsaw operators, heavy equipment operators, drill rig workers).

Biodiversity Management Plan; the Biodiversity Management Plan will require a rapid

survey of flora and fauna (dry season and wet season) at least every five years to monitor

potential changes to the biodiversity profile of the Project. The Biodiversity Management


Detailed Mine Reclamation, and Closure Plan; two years before final closure, a detailed

version of the Mine Reclamation and Closure Plan will be prepared submitted to Guyana

EPA for review and approval. The Detailed Mine Reclamation and Closure Plan will include

specific requirements for monitoring pit lake filling, as well as the completeness of required

demolition or removal actions and the overall effectiveness of the restoration and re-

vegetation of the reclaimed overburden stockpiles, solid waste landfill areas, the TMA, the



vegetation of the reclaimed overburden stockpiles, solid waste landfill areas, the TMA, the

MWP and FWP, the Cuyuni dike areas, roadway ROWs, and other reclaimed areas of the

project. Specific erosional issues associate with closed areas will be monitored (and if

necessary, remediated) in compliance with the Erosion Prevention and Control Plan, which

will be incorporated by reference.

As part of its ESMS, GGI will also develop and maintain a master Environmental, Social, and Health and

Safety Monitoring Plan, which will describe the contents, routine use, and regular update of a

comprehensive, Excel™ based spreadsheet to manage the periodic monitoring requirements defined by

the various management plans described above, as well as other specific monitoring and/or reporting

requirements that may be defined by the IFC PSs and current Guyana EPA Environmental Permit

requirements.

20.3.4 Water Management

The Aurora Gold Project is located in a tropical setting with substantial rainfall, and the management of

water will be among the most significant operational issues encountered at the site. There are sixprimary structures that will be used to manage surface water. These include:

The Cuyuni River dike,

The Fresh Water Pond (FWP),

The Mine Water pond (MWP),


Fresh Water Pond Operation (FWP); the FWP will be constructed to serve as a makeup

water source for the process facility. A dam will be constructed at a crest elevation of 61.0

m to form the pond and retain about 600,000 m3 at the normal high water line (NHWL). The

pond will fill from precipitation over the 140 ha drainage basin. The FWP will discharge into

the MWP through a service spillway.

Mine Water Pond Operation (MWP); water from the open pits underground operations



Mine Water Pond Operation (MWP); water from the open pits, underground operations,

and any other site water not reporting to the TMA will be pumped to the MWP for

clarification and retention.

The MWP will receive water from its 113 ha drainage basin and the open pit mine’s 110 ha drainage

basin. The MWP is also designed to provide sufficient storage capacity for three-day retention of the

runoff volume from a 10-year 24-hour rainfall storm. The pond is sized to receive an average pumping

rate of about 1,000 m3/hr. The average retention time during a normal precipitation year will be about

20 days. Assuming that water quality criteria are met, the pond will be permitted to discharge

continuously through either a low level outlet works or an overflow spillway structure. For flood events

exceeding the 1-in-100-year, 24-hour storm runoff, pond water will be allowed to spill through an

emergency spillway.

Pit dewatering requirements will be reduced by diverting site runoff away from the open pit. The design

concept is to divert runoff away from the open pit in a progressive or staged manner to avoid

accumulating large flows and volumes of surface water near the pit perimeter. This will be achieved bydiverting site runoff that would normally reach the open pit into diversion channels that will eventually

be discharged into the Cuyuni River.

Clean runoff from any areas of the pit which will not be disturbed by stripping or mining activity will be

collected and routed to the environment. In general, runoff will be collected in ditches alongside the site

d d h l d d d h di i h l h h d i l


NamePermitting

AgencyStatus Comments

EnvironmentalPermit

Guyana EPA GrantedSeptember 28,2010

Environmental Permit 20090114-GGIOO wasgranted by the Guyana EPA after reviewing the finalNational ESIA prepared for GGI by GSEC; this

review determined that all EPA, Environmental



Assessment Board (EAB), GGMC, and other agencycomments had been satisfactorily resolved. TheEnvironmental Permit invokes a number of detailedregulatory compliance requirements, as well asrequirements for the lodging of an EnvironmentalBond. GGI has lodged an Environmental Bond infavor of the GGMC, as required by the approvalconditions of the Environmental Permit.

Mining License

(ML)

GGMC Granted

November 18,2011

The ML permits GGI to build and operate the Aurora

Gold Project, and is valid for 20 years withprovisions for extension. It is the first large-scalegold mining ML issued in Guyana since 1991.

Permit to usecyanide

GGMC Pending Before commencing any use of cyanide, GGI mustapply for a special cyanide permit from GGMC,providing information on: the site, design or process, and amount of

cyanide to be used; site characteristics and layout; distance to water bodies; ground water regime; mode of tailings disposal; possible effects on the environment; a simplified description of the activity; and strategies for minimizing the use of cyanide over

the long term.


expected to generate the direct socio-economic effects that often characterize other mining projects.

Most importantly, although the mining concession will require consistent surveillance to identify

potential illegal ASM incursions, there are currently no permanent communities or residences within the

Aurora Gold Project concession that would require physical displacement or resettlement actions.

The area of the project is very remote, but has been impacted by traditional and largely unregulated



p j y , p y g y g

ASM, logging, and hunting activities for well over 100 years. As noted in:

“Technical Report Identifying the Potential Range of Cultural Resources within the Aurora Gol d Mining

Project Area, Guyana” (Plew, 2012) there are no known archaeological sites or areas of significant

cultural interest within the Project concession. However, the Guyana National Trust and Ministry of

Culture have both expressed an interest in any artifacts or items of potential historical, archaeological,

or anthropological interest that may be encountered over the life of the project. Management plans and

standard operating procedures (SOPs) for exploration, construction, and mining operations (and SOPs

for area-specific environmental clearances) will therefore invoke a specific procedure for documenting,

protecting, and reporting chance finds. Implementation of the Community Relations Management Plan

for the Project will also provide the means of detecting and appropriately responding to any changing

stakeholder views with respect to cultural heritage concerns, as well as employment or contracting

opportunities, health and safety, and other social considerations.

The communities considered to be in the Project AOI are described as follows:

The Buckhall Port is the entrance point to the Aurora Gold Project area on the Essequibo

River, and is the location of a concession area used by Guyana Goldfields for the staging and

roadway transportation of materials, equipment, supplies, and employees. The adjacent

community is considered to be in the Aurora Gold Project ’s Direct Area of Influence (DAI).

Aranka Mouth is located approximately 30 km downstream from the Aurora Gold Project


community is Kurutuku, which, due to its remote upriver location, is unlikely to be significantly impacted

by project operations. However, as the Cuyuni River is a significant regional transportation corridor,

Guyana Goldfields does recognize Amerindian groups as potential Aurora Gold Project stakeholders, and

provisions will be made in the Community Relations Management Plan to specifically consider Kurutuku

and other more distant communities as potential recipients of appropriate community investment

programs and other outreach activities Guyana Goldfields currently employs a number of Amerindian



programs and other outreach activities. Guyana Goldfields currently employs a number of Amerindian

workers, and additional hiring opportunities will be made known to Amerindian communities.

The potential for incursions of traditional unregulated or illegal ASM access to Aurora Gold Project areas

by small groups or individual transient miners (or the creation of opportunistic encampments or

informal communities in the area of the Aurora Gold Project) is known to be an area of significant

concern to the Government of Guyana. Such access will therefore be actively discouraged in

collaboration with Government authorities, out of concerns for the health and safety of employees andthe contractor workforce, the transient miners themselves, and other individuals opportunistically

attracted to the Aurora Gold Project because of a perceived potential for economic activity or benefit.

Transient miners and other groups are recognized as potential stakeholders in the Community Relations

Management Plan, however, and will have access to the grievance procedures defined therein.

Transient miners will also have access to potential employment opportunities, subject to the evaluation

of individual qualifications, education, and experience and Guyana Goldfields’ centralized hiring and

controlled workforce transportation practices, as well as expectations for strict compliance with the

code of conduct established for the Guyana Goldfields workforce and contractor staff.

M I N E C L O S U R E 20.6

20.6.1 General Description of Mine Reclamation and Closure Plan

Guyana Goldfields has developed a conceptual Mine Reclamation and Closure Plan for the Project which


The fuel storage tank farm, secondary containments, and fueling station;

The mineral processing facility (including ore sorting, crushing, CIL, and cyanide

detoxification circuits)

The TMA and TMA Diversion Ponds 1 and 2;

The FWP and MWP;L d d h



; Laydown areas and warehouses;

Mechanical and maintenance shops;

Permitted solid waste landfill;

Hazardous waste storage facility area;

Haul and access roads;

Mancamp and administrative buildings; and,

Potable water and septic systems.

20.6.2 Summary of Site Closure and Waste Disposal Strategy

Unless other land uses, mixtures of land use, or other beneficial uses of specific elements of Aurora Gold

Project infrastructure are negotiated with the Government of Guyana and other stakeholders, the

overall goals for Aurora Gold Project decommissioning and closure will be to return the land to a

physically, biologically, and chemically stable and ecologically functional condition that approximates

baseline conditions. Guyana Goldfields is also obliged, as a condition of its Environmental Permit

(Guyana EPA, 2010), to minimize the potential attractiveness of the decommissioned site for illegal or

uncontrolled ASM activities. Progressive closure options will therefore be sought, wherever possible in

the construction and operational phases of mine life, in an effort to minimize the potential for

subsidence and erosion damage, to enhance biodiversity and the restoration of natural habitats, and to

minimize the potential attractiveness of the site. These options will include:


Selective breaching, modification, and re-vegetation of the Cuyuni River dikes, and selective

breaching, regrading, and re-vegetation of the FWP, MWP, TMA, and TMA Diversion Pond

embankments, as indicated in Figure 20-4;

Placement of soil cover and revegetation of the dewatered beach areas of the TMA;

Construction of effluent settling, dilution, and/or polishing ponds for the FWP, MWP, TMA,and TMA Diversion Ponds if necessary to ensure consistency of discharged water quality



g p g pand TMA Diversion Ponds if necessary to ensure consistency of discharged water quality

with respect to the (IFC, 2007) effluent standards;

Development of interconnected pit lakes and a stable natural drainage channel to the

Cuyuni River (with a security berm and warning signs established throughout the period of

pit lake infilling); and

Controlled closure and abandonment of monitoring wells and piezometers at the end of

post-closure monitoring.

It should be noted that Guyana does not currently have any permitted hazardous waste disposal

facilities; however, one facility is planned and this facility will be available for use at the time of mine

closure. It is also noted that Guyana does not currently have any significant metals or other waste

materials recycling capabilities. Guyana Goldfields will monitor for the development of such capabilities

over the years of mine operation, and will update the Mine Reclamation and Closure Plan accordingly if

viable recycling sources are identified for any of the waste types generated in site decommissioning and

closure. However, for the purposes of this Technical Report and the initial version of the Mine

Reclamation and Closure Plan it is conservatively assumed that no recycling facilities will be available.With these assumptions in mind, decommissioning wastes will be managed as follows:

Major equipment items are not expected to have any significant resale value at closure are

limited to the crushers, the ferry and barge from Tapir Crossing, the TMA reclaim barge,

high-value gold room equipment, and a number of the larger pumps, will be removed at


Cyanide facility equipment, tanks, and piping systems will be flushed, with rinseate routed

to the detoxification plant prior to disposal in the TMA, after which the rinsed detoxification

plant and the tailings pipelines will also be demolished and disposed of in the underground

mine or inert waste cell in the waste rock stockpile.

Residual hazardous materials (e.g., unused reagents, fuel, lubricants, paints, insecticides, reagents, or



explosives) will be returned to suppliers for credit or otherwise sold to properly licensed or reputable

dealers, and strictly for the purposes intended by the manufacturer. Any residual hazardous wastes will

be accumulated in the onsite hazardous waste storage facility prior to being routed to an approved

offsite hazardous waste landfill; residual medical wastes will be routed to an approved medical waste

incinerator in Georgetown.

20.6.3 Cost Estimate

A preliminary estimate of the conceptual costs of closure for the Aurora Gold Project at the end of mine

life is presented in Table 20-3.

Table 20–3: Estimated End of Mine Life Closure Costs

PROJECT TASK AREA COST

MINE AND WASTE ROCK 424,539

Cuyuni River Dike 1 161,985Cuyuni River Dike 2 97,191

Underground Mine / Facilities 52,000

Open Pit 113,363

Waste Rock Stockpiles 0

PROCESS FACILITY 518 082


PROJECT TASK AREA COST

TOTAL PROJECTED DIRECT COSTS $3,823,058

INDIRECT COSTS

Engineering / Design (5% of directs) 191,153

Contingency (15% of directs) 573,459

http://c/Users/gmills/Desktop/GGI%20%20Decomissioning%20&%20Closure%20Pan%20Proposal/GGI-CCM%20V1.0%20(12-07-2012)%20BASE.xlsx%23'3.1%208100%20MINE%20&%20WASTE%20ROCK'!A1









http://c/Users/gmills/Desktop/GGI%20%20Decomissioning%20&%20Closure%20Pan%20Proposal/GGI-CCM%20V1.0%20(12-07-2012)%20BASE.xlsx%23'3.2%208300%20PROCESS%20FACILITY'!A1

http://c/Users/gmills/Desktop/GGI%20%20Decomissioning%20&%20Closure%20Pan%20Proposal/GGI-CCM%20V1.0%20(12-07-2012)%20BASE.xlsx%23'3.2%208300%20PROCESS%20FACILITY'!A1









Unlisted Items (10% of directs) 382,306

Mobilization / Demobilization / Preparatory Work (10% of directs) 382,306

Construction Engineering / Construction Management 212,846

Monitoring and Maintenance (2 years) 478,514

Equipment Salvage (345,000)

Retrenchment 3,201,885

Permitting (1% of directs) 38,231TOTAL PROJECTED INDIRECT COSTS $5,115,700

TOTAL PROJECTED PROJECT COSTS (rounded) $9,000,000

These assumptions are integrated into the initial closure cost estimate presented in Table 20-3. In

addition, as a condition of Guyana Goldfields’ current Environmental Permit (Guyana EPA, 2010), two

years prior to cessation of all mining operations and the decommissioning and demolition of ore

processing and support facilities, a Detailed Mine Reclamation and Closure Plan (DRMCP) prepared, withappropriate input from affected stakeholders, and presented to the Guyanese EPA for review and

approval. Any subsequent updates of this DMRCP shall be issued on a schedule to be negotiated with

the Guyana EPA.

20.6.4 Post-Closure Monitoring


Continued monitoring for the colonization of reclaimed areas by native fauna, in accordance

with the Biodiversity Management Plan.




2 1 . 0 C A P I T A L A N D O P E R AT I N G C O S T S

Capital and operating costs results summarized in this section are based upon work performed by

various third party engineers and consultants and representatives from Guyana Goldfields. SRK prepared



p y g p y p p

capital and operating costs associated with open pit and underground mining. Tetra Tech updated

original work completed by Ausenco with respect to process and infrastructure. Owner’s costs were

estimated by Tetra Tech and Environ with assistance from Guyana Goldfields.

Cost estimates are based upon first principle estimates and are presented in December 2012 US dollars.

No escalation is applied to capital or operating costs. Project cost estimates and economics are prepared

on a quarterly basis for the calendar years 2013 through to 2017.

Capital and operating costs tables presented in this report require subsequent calculations to derive

subtotals, totals, and weighted averages. Such calculations inherently involve a degree of rounding.

Where these occur they are not considered to be material.

Based upon design criteria presented in this report, the level of accuracy of the estimate is considered

±15%.

C A P I T A L C O S T S 21.1

Life-of-Mine (LoM) Project capital is summarized in Table 21-1. Initial capital costs are estimated at $205

million. Expansion capital includes; $93 million for underground mine development (ramp, ventilation

raises, etc.) as well as mining equipment, and $27 million for expansion of the process and power plant.


Cost estimates are supported by budget quotations from equipment vendors, including allowances for

freight and assembly.

Table 21–2: Open Pit Capital ( 000s)

1

Capital Costs

Initial

(2013 –2014)

Expansion

(2015-2017)

Sustaining

(2018 –2031) Total

Mi P d ti $2 261 $8 822 $6 914 $17 997



Mine Production $2,261 $8,822 $6,914 $17,997

Mine Support $4,038 $0 $1,854 $5,892

Blasting Facilities $106 $722 $0 $828

Ancillary Equipment $2,910 $13 $13 $2,936

Other $3,226 $1,438 $1,611 $6,275

Total1 $12,541 $10,995 $10,392 $33,928

1. Includes contingency (5% mobile equipment, 10% other). Addition differences due to rounding

The following describes the cost categories shown in the table.

Mine production equipment represent units acquired for the drilling, loading and hauling.

Mine Support equipment include; tractors, rubber-tired dozers, graders and water trucks.

Blasting facilities capital include; an MMU truck, silos, magazines and general support

equipment.

Ancillary mobile equipment cost is for mining related service equipment and smaller support

units. Other capital costs include estimates for freight, technical services, pit dewatering and haul

road construction.

Contingencies of 5% are used on mobile equipment and 10% on all other items.

21 1 2 Underground Mine


Table 21–3: Underground Capital (US 000s)

Capital CostsInitial

(2015-2017)Sustaining

(2018-2031)Total

Direct CostsPortal and Contractor Development $43,756 $10,343 $54,099

Contractor Diamond Drilling $0 $21,091 $21,091Owner Capital Development $0 $63,339 $63,339M bil E i Fl $9 477 $124 512 $133 989



Mobile Equipment Fleet $9,477 $124,512 $133,989Ventilation and Cooling $8,574 $15,051 $23,625Mine Services $5,298 $18,707 $24,005Safety and Egress Ladder Way $1,978 $12,047 $14,025Material Handling Infrastructure $0 $5,746 $5,746Technical Services $649 $1,395 $2,044

Direct Costs $69,732 $272,230 $341,963Indirect & ContingencyIndirect Costs and EPCM $12,420 $2,604 $15,024

Contingency $10,460 $40,835 $51,294Indirect Costs $22,880 $43,438 $66,318

Total $92,612 $315,669 $408,281


Work on the underground portal and decline development will commence in Q4 2015. The Aurora Gold

Project total underground capital cost estimate is $408 million, comprised of $93 million initial capital

and $316 million sustaining capital. A 15% contingency was applied to direct initial capital costs. Capital

costs include estimates for:

Portal construction and production decline;

Off decline development;

Lateral development;

Large diameter raise boring;




Table 21–5: Owner & Closure Capital ( 000s)

CapitalInitial

(2013-2014)Expansion

(2015-2017)Sustaining

(2018-2031) Total

Owner's Costs $16,545 $1,811 $0 $18,356

Mine Closure $0 $0 $9,000 $9,000

Total $16,545 $1,811 $9,000 $27,356Addition differences d e to ro nding




Owner costs include construction camp costs including catering, inland freight from Buckhall Port to

project site, upgrades at Tapir Crossing, ocean insurance and training. As discussed earlier, mine closure

costs are estimated to be $9 million.

OP E R A T I N G C O S T S 21.2

LoM operating costs are summarized in Table 21-6. LoM operating costs are estimated at $1.4 billion, or

$34.95/t-milled. Open pit mining will average $2.42/t-moved ($13.68/t-ore). Underground mining will

be $19.28/t-ore over the LoM. Process and G&A costs are estimated at $13.78/t-milled and $3.83/t-

milled, respectively.

Table 21

–

6: LoM Operating Costs

Cost ItemLoM Cost

($000s)

Unit Cost

$/t-moved

Unit Cost

$/t-ore

Unit Cost

$/t-milledOpen Pit Mining $186,999 $2.42 $13.68 -

Underground Mining $498,435 - $19.28 -

Processing $544,551 - - $13.78

G&A $151,225 - - $3.83

Operating Costs $1,381,209 $34.95


Table 21–7: Open Pit Operating Costs (LoM)

Cost ItemLoM Cost

($000s)Unit Cost

$/t-totalUnit Cost

$/t-ore

Mine Ops - General $14,560 $0.188 $1.065

Drilling $9,651 $0.125 $0.706

Blasting $40,983 $0.529 $2.997

L di $18 098 $0 234 $1 324



Loading $18,098 $0.234 $1.324

Hauling $39,769 $0.514 $2.908

Tractor $21,566 $0.279 $1.577

Grader $9,805 $0.127 $0.717

RTD $6,369 $0.082 $0.466

Water Truck $1,673 $0.022 $0.122

Pickup Truck $6,188 $0.080 $0.453

Haul Roads $2,069 $0.027 $0.151Pit Dewatering $5,028 $0.065 $0.368

Reclamation $11,239 $0.145 $0.822

OP Operating Cost $186,999 $2.416 $13.676


21.2.2 Underground Mine

LoM underground mining costs include $76 million for ore development and $422 million in direct costs.

Combined, underground operating cost is estimated to be $19.28/t-ore. LoM costs are summarized in

Table 21-8.

Ore development costs to develop 5m height x 5.5m width headings will average $3,020/m, ranging


Table 21–8: Underground Operating Costs (LoM)

Cost ItemLoM Cost

($000s)Development

(meters)Unit Cost

$/mUnit Cost

$/t-ore

Development

5m x 5.5m headings $76,422 25,308 $3,020 $2.96

UG Operating

Mining



Mining

LH drilling $22,360 - - $0.86

ITH V30 slot drilling $5,370 - - $0.21

Blasting $40,084 - - $1.55

Bogging $35,458 - - $1.37

Hauling $103,061 - - $3.99

Subtotal $206,333 - - $7.98

Utilities & Labor Auxiliary equipment $19,320 - - $0.75

Power $111,962 - - $4.33

Labor $78,455 - - $3.03

Dewatering $5,203 - - $0.20

Ventilation $741 - - $0.03

Subtotal $215,681 - - $8.34

Mining $422,013 - - $16.32

UG Operating Cost $498,435 - - $19.28Addition differences due to rounding

21.2.3 CIL Process Plant


Electric Power

Electric power will be generated on site to service mining, processing and all site service power

requirements using reciprocating engine generators burning No 4. Fuel Oil. No. 4 Fuel Oil is a mix of No.

2 oil (diesel fuel) and No. 6 oil (HFO/bunker oil). No. 4 oil is selected as a compromise between higher

cost diesel fuel and HFO which is difficult to transport and handle.



Power costs are estimated to be $0.228/kWh. The estimate is based upon a quoted Fuel Oil cost of

$0.89/liter ($0.208/kWh) delivered to Buckhall Port and an O&M cost of $0.020/kWh for operating the

generators. The cost of transporting fuel to site from Buckhall Port is included in the G&A cost estimate.

CIL power requirements are summarized in Table 21-10 for 5,000 t/d and 10,000 t/d operations.

Table 21–10: CIL Plant Power Requirements

Facility AreaRequirement

@5,000 t/d (kWh/t)Requirement

@10,000 t/d (kWh/t)

Primary Crushing 0.6 0.4

Secondary Crushing 0.9 0.6

Tertiary Crushing 2.2 2.2

Conveying 0.6 0.5

Grinding 20.7 20.7

Pre-Leach Thickener 0.6 0.3

Carbon-in-Leach 1.5 1.4Desorption 0.2 0.1

Gold Room 1.3 0.7

Reagents 0.5 0.3

Detox./Tails Transport 1.2 0.8

Service/Utilities 1 7 1 0


Nitric Acid 0.08 kg/t $765.64 per tonne $0.061Diesel 0.38 L/t $1.09 per Liter $0.409Borax 0.40 g/t $1,492.94 per tonne $0.001Silica 0.15 kg/kg $655.00 per tonne $0.000Soda Ash 0.10 kg/kg $706.15 per tonne $0.000Potassium Nitrate 0.03 kg/kg $2,044.09 per tonne $0.000

Unit Cost $4.083Addition differences due to rounding




Consumables

Consumable costs will average $1.61/t-milled over the LoM, varying from$1.45/t-milled at 5,000 t/d and

$1.71/t-milled at 10,000 t/d as shown in Table 21-12.

Table 21–12: Consumable Costs

Facility AreaCons

@ 5kt/dCons

@ 10kt/d

ConsumableCost

(US$)

Unit Cost@ 5kt/d

$/t-milled

Unit Cost@ 10kt/d$/t-milled

Jaw Crusher Liner 5 10 sets/yr $15,885 $0.045 $0.091Sec. Cone Crusher Liner 5 10 sets/yr $16,947 $0.048 $0.097Tert. Cone Crusher Liner 5 10 sets/yr $16,947 $0.048 $0.097Ball Mill Liner 1 1 sets/yr $430,100 $0.123 $0.246

Ball Mill Balls 1 1 kg/t $1,345 $1.130 $1.130Slaking Mill Liner 1 1 sets/yr $34,040 $0.010 $0.010Slaking Mill Balls 0.03 0.03 kg/t $1,357 $0.041 $0.041

Unit Cost $1.445 $1.710



Table 21–13: General & Administrative Costs

Cost ItemLoM Cost


Buckhall $63,459 $1.606

Tapir Crossing $6,935 $0.175

Roads $13,164 $0.333Clearing $686 $0 017



Clearing $686 $0.017

Aurora Camp $66,981 $1.695

G&A Costs $151,225 $3.826


Operations at Buckhall Port are primarily logistics focused. Costs address handling equipment, supplies,

fuel and reagents arriving by ship or barge, clearing customs, and transportation to the project site. On

average, about 66 personnel will be assigned to this area.

Tapir crossing operations will support barging equipment across the Cuyuni River. This function requires

the operation and maintenance of the barge, and security. Twenty personnel staff will be assigned to

this operation.

Roads operations are assigned to maintaining the mine access road to assure continuing supply service

to the operation. Thirteen personnel will be assigned to this task.

Clearing operations will occur during pre-production and the first 3 years of production, ending Q2 2017.

Twelve personnel will be involved in clearing trees in support of site development activity.

On average, 151 personnel will be employed performing general and administrative functions.


2 2 . 0 E C O N O M I C A N A L Y S I S

Technical-Economic costs results summarized in this section are based upon work performed by Tetra

T h d G G ldfi ld ’ l



Tech and Guyana Goldfields’ personnel.

Economics are presented in December 2012 US dollars. No escalation is applied to capital or operating

costs. Project cost estimates and economics are prepared on a quarterly basis for the calendar years

2013 through 2017.

Based upon design criteria presented in this report, the level of accuracy of the estimate is considered

±15%.

The Technical-Economic Model (TEM) is presented in Figure 22-2.

Technical economic tables and figures presented in this report require subsequent calculations to derive

subtotals, totals, and weighted averages. Such calculations inherently involve a degree of rounding.

Where these occur they are not considered to be material.

PR I N C I P A L A S S U M P T I O N S 22.122.1.1 Technical Parameters

Parameters used in the TEM are shown in Table 22-1. These parameters are based upon current market

conditions, vendor quotes, design criteria developed by Guyana Goldfields personnel, and benchmarks

against similar existing projects Government royalty and income tax assumptions reflect the Mining


Refiners currently pay for over 99.9% of gold in doré and charge approximately $0.35/oz-Au.

Transportation costs from the site by helicopter to a refinery pick up point, then by the refiner to its

refinery are presented on a per ounce doré basis at $2.70/oz.

Table 22–1: TEM Principal Assumptions

Description Parameter



General Assumptions:

Pre-Production Period 24 months

Mine Life 17 years

Operating Days 350 day/yr

Initial Capacity 5,000 t/d

Expanded Capacity 10,000 t/d

Market Assumptions:

Gold Price $1,300 per Au-oz

Refinery Terms:

Payable Metal 99.925% Au in doré

Refinery Charge $0.350 per payable oz

Freight & Insurance $2.700 per oz doré

Financial Assumptions:

Government Royalty 8% NSR

Income Tax 30%

Cost of Capital 5%

Gearing None

Technical Assumptions:

Fuel Oil Price $0.89 per Liter

Diesel Fuel $0.95 Per Liter

Power Cost $0 228 per kWh


Table 22–2: Production Summary

ReservesOre(kt)

Grade(g/t)

ContainedAu (koz)

Open Pit

Saprolite 4,790 1.72 266

OP Fresh Rock 8,057 3.09 801Stockpile 826 2 02 54



Stockpile 826 2.02 54

OP Total 13,673 2.55 1,120

Underground 25,851 2.84 2,357

Total 39,524 2.74 3,477


C A S H F L O W 22.2A LoM cash flow analysis is provided in Figure 22-2.

E C O N O M I C R E S U L T S 22.3

Economic results are summarized in Table 22-3. The analysis suggests the following conclusions

assuming no gearing:

Mine Life: 17 years

Pre-Tax NPV5%: $1,119 million, IRR: 44%

Post-Tax NPV5%: $800 million, IRR: 38%

Payback (Post-Tax): 40 months

Corporate Income Taxes Paid: $509 million


Table 22–3: Technical-Economic Results

DescriptionLoM Cost


Unit Cost$/oz-Au

Gross Revenue $4,277,922 $108.24 $1,300.00

Refining ($13,238) ($0.33) ($4.03)

NSR $4,264,684 $107.90 $1,296.95

Royalty ($341 175) ($8 63) ($103 76)



Royalty ($341,175) ($8.63) ($103.76)

Net Revenue $3,923,509 $99.27 $1,193.19

Operating Costs

Mining $685,434 $17.34 $208.45

Processing $544,551 $13.78 $165.61

G&A $151,225 $3.83 $45.99

Operating Costs $1,381,209 $34.95 $420.04

Operating Profit $2,542,300 $64.32 $773.15Capital Costs

Capitalized Costs $138,529 - -

Mining $303,680 - -

Process & Infra. $244,583 - -

Owner's Costs $27,356 - -

Total Capital $714,148 - -

Pre-Tax Cash Flow $1,828,152 - -

NPV5% $1,118,843 - - IRR 44% - -

Payback (months) 38 - -

Peak Funding ($162,889) - -

Post-Tax Cash Flow $1,319,247 - -

NPV5% $799,720 - -


Duty and value added tax exemptions on all imports of equipment and materials for all

continuing operations at the Aurora Gold Project, including the construction and operation

of a planned port facility, road and power improvements and the construction and

operation of the mine.

The Mining Licence is the Company's permit to build and operate mining facilities at the Aurora GoldProject and is valid for an initial 20-year term with provisions for extension on application by the



Project and is valid for an initial 20 year term with provisions for extension on application by the

Company.

This royalty is assessed on the gross value sales of gold less doré transportation and refining costs. Over

the LoM, royalties will total $341 million or $104/oz payable gold at a price of gold of $1,300/oz. Royalty

accounts for approximately 20% of cash costs.

22.4.2 TaxesUnder the Agreements signed effective November 18, 2011, the company shall pay income tax pursuant

to the Income Tax Act in Guyana generally, provided that the corporate tax payable shall be the lesser of

(i) 30% and (ii) the prevailing rate in Guyana from time to time during the Agreements’ term. The tax

rate currently payable is 30% of chargeable profit.

The amount of income tax payable shall be calculated in accordance with the tax rules as they apply as

of the effective date of the Agreements, to the exclusion of such rules less favorable to the Company as

may be enacted in the future. However should any subsequent amendment to the tax rules result in the

enactment of any rules regarding the calculation of taxes which are more favorable to the Company

(than those in effect on November 18, 2011), the Company shall be entitled to invoke such amended

rules for the purposes of calculating its tax liability.

In calculating the Company’s taxable income the Company shall have the right to amortize:






January 2013 290

Figure 22 –2 Life of Mine Cash Flow Projection


2 3 . 0 A D J AC E N T P R O P E R T I E S



There are no adjacent properties that are considered relevant to this technical report.


2 4 . 0 O T H E R R E L E V A N T D ATA A N D

I N F O R M A T I O N



PR O J E C T R I S K S – T E C H N I C A L 24.1

In this section, potential technical risks are discussed. This includes risks stemming from uncertainties in

geotechnical and hydrogeological characterization and potential risks related to the mining method.

24.1.1 Geotechnical Risks

Geotechnical risks are present for any greenfield project where no previous mining history and

experiences with the rock mass behavior and response to mining conditions have been tested. Those

risks are stemming from the limitations of drill hole based geotechnical and structural data and residual

uncertainty in selected base case parameters. In such cases, it is important to assess potential ranges of

outcomes and apply both empirical and numerical methods of rock mass strength and stress evaluation

and conduct sensitivity analyses to understand the consequences of potential deviation from the base

case. SRK in collaboration with Itasca followed such process.

As part of the risk mitigation strategy, it was important to develop a design that has flexibility to address

potential unexpected situations; such as hidden and undetected faults, or deviation from the ore body

geometry etc., without major interruption to the mining process. The second important risk mitigation

h d l d f h d


sensitivity analyses indicated that it is unlikely that stress changes within a reasonable range

of magnitude or direction could result in uncontrollable collapse of the stope walls. It is also

important to understand that even if such an unlikely situation would occur, the monitoring

program would provide an early warning of changing conditions and mitigation measures;

such as waste backfilling of the SLR stope excavation could be implemented and SLR mining



such as waste backfilling of the SLR stope excavation could be implemented and SLR mining

would change to sublevel “caving.”

The second residual risk is the presence of unknown large scale structures. We have

concluded that with the amount of drilling and hydrogeological testing it is highly unlikely

that any unknown structures would cause a major impact on the mining plan. Again,

mitigation measures in terms of grouting and increasing pumping or storage capacity would

be available.

The open pit design criterion is currently based on the previous geotechnical domains (AMEC, 2009). The

updated geological and geomechanical domains include the sericite shear band, which has a distinct

strong foliation in which rock breakage occurs. These shear zones will need to be incorporated into an

updated open pit design. Although the strongly foliated rock may present some risk in terms of toppling

behavior within the north pit wall at Rory’s Knoll, the sub vertical nature and limited extent in the pit

wall should not have a material impact on the operation.

24.1.2 Open Pit Mine Risks

The proposed mining operation is located in a region that receives significant tropical rainstorms that


the decline and there is provision to increase pumping capacity. Although this would increase the

operating cost, it would not be a fatal flaw in terms of the mine design.

External mudrush risk exists for the underground mine due to the heavy rainfall and the potential for

generating fines and clays from the overlaying saprolite material This risk will be mitigated by partial



generating fines and clays from the overlaying saprolite material. This risk will be mitigated by partial

pre-stripping of saprolites as part of the open pit mining and by implementation of proper dewatering

and water diversion programs, such as perimeter drainage, collection sumps, etc.

Timely supply of expatriate and skilled local personnel has the potential to be a very significant risk to

the success of the project. The ability to adequately train local un-skilled labour to the required level is

also a key factor for the underground mine. To mitigate this risk, it is assumed that in the years the minewill be developed using an experienced underground contractor, a comprehensive training program is

introduced.

24.1.4 Mineral Processing

A full risk assessment of the transportation of reagents and consumables to site should be conducted to

determine any logistics issues given the plant site location.

Whilst there is no shortage of water on-site, a clean source of fresh water has yet to be finalized, and

should be identified early in the next phase of the project.

24.1.5 Infrastructure Risks

Use of trained lo al onstr tion orkfor e alon ith ood onstr tion mana ement has been


Re-optimize the pits as the FS operating cost is lower than the preliminary operating cost

used in the study.

Additional pit designs work to include the smaller satellite pits which were excluded from

this study.

Analyzing the production schedule for opportunities to reduce the strip ratio late in the



Analyzing the production schedule for opportunities to reduce the strip ratio late in the

mine life and reduce the expansion capital required.

Phasing out more of the expatriate staff over the mine life to reduce labour costs.

Optimization of the pit slope angles to reduce stripping ratio.

Optimization of drilling and blasting designs to improve productivities and mill throughput.

Investigate the value of optimizing each of the satellite pits as an open pit / underground

crossover.

24.2.2 Underground Mine Opportunities

The underground plan presented has several remaining opportunities to investigate:

Evaluate implementing an overdraw strategy at the end of the mine life to mine the low

grade mineralized material left in the SLR stope excavation.

Analyze the possibility to economically mine mineralized material below the other pits.

Analyze the possibility to mine below the 970 mbsl using the SLR mining method.

Evaluate alternative lower cost haulage systems (i.e. Rail-Veyor material haulage

technology) which have the potential to lower the underground capital and operating costs.

Mine deeper if walls are more stable than expected and/or implement a backfill mining

method


2 5 . 0 I N T E R P R E T AT I O N A N D C O N C L U S I O N S



G E O L O G Y A N D R E S O U R C E S 25.1

Exploration work is professionally managed and field procedures generally meet accepted

industry best practices. SRK is of the opinion that the exploration data are sufficiently

reliable to interpret with confidence the boundaries of the gold mineralization and support

evaluation and classification of mineral resources in accordance with generally accepted

CIM “Estimation of Mineral Resource and Mineral Reserve Best Practices” and CIM“Definition Standards for Mineral Resources and Mineral Reserves” guidelines;

The bulk of the mineral resources are located in Rory’s Knoll, which represents 69% of the

total reported Measured and Indicated mineral resources and 73% of the reported Inferred

mineral resources;

The Aurora gold deposit contains a significant mineral resource estimated at 6.54 million

ounces of gold in the Measured and Indicated categories with an additional 1.82 million

ounces of gold in the Inferred category. SRK notes that the mineral resources occupy a small

footprint on the prospecting license.

OP E N P I T M I N E C O N C L U S I O N S 25.2

The near surface mineralization at the Aurora Gold Project is amenable to conventional

loader/truck mining methods utilizing 7 7m3 front end loaders and 43 5 tonne class


This mining method has precedent in South Africa and Canada, where two pipe type

deposits are being mined using this method.

Financial modeling of the underground has determined that underground mining is

economically viable and supports Probable Reserves of 25.8 Mt of ore grading 2.84 g/t gold.





commencing in the fourth quarter of 2015, with underground commercial production

beginning in early 2018.

The underground mine will feed the mill at a nominal rate of 1.9M tonnes per year.

The underground mine will require US$92.6M of initial pre-production capital and

US$315.7M of sustaining capital.

The average operating cost of the underground mine is US$19.28 per ore tonne mined. A comprehensive underground geotechnical instrumentation and monitoring program has

to be implemented to mitigate potential risk of larger than expected stope wall failures

(refer to the technical risks section).

A comprehensive training program will be required to train local labour for the underground

mine.

M I N E R A L P R O C E S S I N G 25.4 Testwork indicated that the Aurora ore could be processed in a conventional circuit

incorporating grinding, CIL, carbon desorption and eluate electrowinning.

Cyanide tailings can be detoxified using an Air/SO2 method;

Ores are amenable to cyanide leaching with gold recoveries consistently in the range of 90


E N V I R O N M E N T A L A N D S O C I A L C O N C L U S I O N S 25.6

The project’s area of influence (AOI) has been significantly impacted by historical artisanal

and small-scale mining (ASM), logging, and hunting, for well over a hundred years; Large fauna that are otherwise common in pristine habitats along similar types of rivers in



this area of South America are absent or rare in the project AOI, and may be viewed as a key

indicator of significant historical human impact;

With very few exceptions, rare, threatened, or endangered species have not been observed

in the area of the project;

There are no formal or established communities or settlements in the immediate vicinity of

the Aurora site, and the project is not expected to generate direct socio-economic effects;

Access by traditional unregulated or illegal ASM to the project vicinity by transient miners is

of concern;

There is no evidence of indigenous hunting activity within the proposed mining area;

Results of geochemical testing to date indicate that project overburden and waste rock has

very low acid rock drainage (ARD)/metals leaching potential;

The management of tailings and mining-impacted water is designed for high-precipitation

climatic conditions; and

The project will develop and implement a comprehensive Environmental and SocialManagement System (ESMS).


2 6 . 0 R E C O M M E N D A T I O N S

G E O L O G Y A N D R E S O U R C E S26 1



G E O L O G Y A N D R E S O U R C E S 26.1

SRK considers that the mineral resource model documented herein is sufficiently reliable to support

engineering and design studies to evaluate the viability of a mining project at a feasibility level. SRK do

however recommend a two component work program to further enhance the resource.

Infill with reverse circulation and core drilling to further expand the mineral resources and

improve resource classification; and

Geological studies aimed at improving the understanding of the geological and structural

setting of the deposit.

SRK consider that additional drilling is warranted to infill areas currently classified as Inferred and to

target areas with the potential for lateral and depth extensions of currently modeled gold

mineralization. Further condemnation drilling is also recommended to support mine infrastructure

design.

Considerable advances in the understanding of the geological and structural controls on the distribution

of the gold mineralization were made in recent years. SRK recommends that further geological studies

be initiated in particularly the satellite deposits to build on existing knowledge and improve the

confidence in the interpretation of the boundaries of the gold mineralization, understand its distribution


Development of a detailed open pit and underground instrumentation and monitoring

program.

Conduction of in-situ stress testing during the pre-development and production

underground phases.





standard operating procedure for such events.

Estimated budget for the open pit evaluation is US$50,000. Estimated budget for the underground

drilling, characterization, and modeling is US$150,000.

An estimate is not given at this time for stress testing, mudrush evaluation, and portal investigation.

H Y D R O G E O L O G I C A L R E C O M M E N D A T I O N S 26.3

It is recommended that further detailed hydrogeological studies are completed to confirm expected

hydraulic conductivity of the zone between the Cuyuni River and the mine area and to test the hydraulic

conductivity of the shear zones. The studies would incorporate the following:

Prior to pump tests being conducted, multi-level piezometers are to be located and installed

at 5-6 locations and will be used in analyzing the groundwater conditions at the site. Pump testing to be conducted in the alluvium/saprolite unit, the weathered bedrock and in

the shear zones.

Recalibration of the groundwater flow model to the new pumping test results and simulate

inflow rates to the open pit and underground mines.


U N D E R G R O U N D M I N E 26.5

It is recommended that further detailed engineering studies incorporate the following:

Performance of a tradeoff study to evaluate the Rail-Veyor material haulage system as an

alternative to conventional truck haulage



alternative to conventional truck haulage.

Performance of a scoping study to evaluate an amendable and economical underground

mining method to extract additional Mineral Resources below the satellite pits.

Performance of a scoping study to evaluate the optimum underground mining depth by the

SLR mining methods.

Estimated budget for this work is US$300,000.

M I N E R A L P R O C E S S I N G 26.6

A trade-off study comparing a 3-stage crush plus ball circuit (3CB) against a high pressure

grinding roll (HPGR) coupled with a tower mill should be undertaken. Such a circuit is likely

more capital intensive, but will result in lower operating cost due to lower power

consumption. Estimated budget for this work is $30,000;

Detailed process facility engineering will be required to prepare construction-readydocuments and to finalize construction cost estimates. Estimated budget for this work is

$4,000,000;

It is suggested that new testing and investigation be performed should new rock/ore types

be encountered during exploration. Estimated budget for this work is $25,000;


B U D G E T 26.8

Estimated budgets have been included with the individual recommendations.




2 7 . 0 R E F E R E N C E S

Bracewell S 1949 The Geology of the Aurora District Cuyuni River British Guiana: Bulletin No



Bracewell, S., 1949, The Geology of the Aurora District Cuyuni River, British Guiana: Bulletin No.

21, British Guiana Geological Survey.

Bluhm Burton Engineering (Pty) Ltd, November 2011: Aurora Project Feasibility Ventilation and

Refrigeration

Cargill, D.G., 2005, NI 43-101 Report on the Aurora Project Guyana 2005, technical report

prepared by Cargill Consulting Geologists Ltd. for Guyana Goldfields Inc., effective date

30 December, 2005.

Cargill, D.G. and Gow, N.N., 2003, Report on the Aurora Project, Guyana: technical report

prepared for Guyana Goldfields Inc., effective date 30 August, 2003.

Gibbs, A.K. and Barron, C.N., 1993, The Geology of the Guiana Shield: Oxford University Press,

247 p.

Golder Paste Technology Ltd, March 9, 2012: Feasibility Report for Paste Backfill

AMEC Americas Ltd, October 3, 2011: Phase 1 Groundwater Flow Model Update

AMEC Americas Limited, June 2, 2009: Aurora Gold Project, NI 43-101 Technical Report on


Kipfel, p., 2005b, Petrography Update: Unpublished Report for Guyana Goldfields

McConnell, R.B. and Williams, E., 1969. Distribution and Provisional Correlation of the

Precambrian of the Guiana Shield. In Proceedings of the Eighth Guiana GeologicalConference, Georgetown Guyana., pp. I-2 to I-22 + Maps. Government of Guyana



Geological Survey Department.

Montejo, M., Kilpatrick, R, Smith, J, Magumbe, L., Zhang, B., and Szymanski, M., 2009, Aurora

Gold Project Guyana, South America, NI43-101 Technical Report on Updated Preliminary

Assessment, Technical Report prepared by AMEC for Guyana Goldfields Inc., effective

date June 2, 2009.

Mukhopadhyay, M.K., 2007, Technical Report: A Mineral Resource Estimate for the Rory’s Knoll,

East Walcott and Aleck Hill Zones, Aurora Gold Property, Guyana, technical report

prepared by Micon International for Guyana Goldfields Inc., effective date 21

November, 2007.

Myers, P.,2008 Guyana Goldfields: Aurora Project Preliminary Assessment, Project No.7178:

technical report prepared by Snowden Associates for Guyana Goldfields Inc., effective

date 14 August, 2008.

Voicu, G., Bardoux, M., Jebrak, M., and Crepau, R., 1999, Structural, Mineralogical and

Geochemical Studies of the Paleoproterozoic Omai Gold Deposit, Guyana: Economic

Geology, 94, pp. 1277-1304.


GOLDFIENDS INC., Lakefield, Ontario; SGS Lakefield Research Limited, Project 12088-002

- Final Report.

Hendry, L., Trang, C. And Jackman, R., 2010: An Investigation of THE CHARACTERISATION OFSAMPLES FROM THE AURORA GOLD PROJECT - FEASIBILITY PHASE prepared for GUYANA



GOLDFIELDS INC., Lakefield, Ontario; SGS Canada Inc., Project 12088-005 - Final Report

#1.

Mezei, A. and Ashbury, M., 2010: An Investigation into the LIQUID-SOLID SEPARATION

RESPONSE OF AURORA GOLD PROJECT SAMPLES - FEASIBILITY PHASE Prepared for

GUYANA GOLDFIELDS INC., Lakefield, Ontario; SGS Canada Inc., Project 12088-005 -

Report #2.

Verret, F. and McKen, A., 2010: An Investigation into GRINDABILITY TESTING OF SAMPLES

FROM THE AURORA PROJECT prepared for AMEC on behalf of GUYANA GOLDFIELDS

INC., Lakefield, Ontario; SGS Canada Inc., Project 12088-005 - Grindability Report.

Chenje, T. and Radziszewski, P., 2010: Aurora Project Grinding Media Wear Prediction,

Montreal, Quebec; CDLabServices, Lab Report.

Contract Support Services, 2010: Drop Weight Test Report on Three Samples from Aurora, Red

Bluff, California; Contract Support Services, SGS Job No. 12088-005.

Cole, G., McInnis, C. and Couture, J-F., 2011: Mineral Resource Evaluation, Aurora Gold Project,

Guyana Report Prepared for Guyana Goldfields Inc Toronto Ontario; SRK Consulting


GSEC, 2007: Final Report Environmental and Social Baseline Aurora Mining Concession for

Guyana Goldfields, technical report prepared by Ground Structures Engineering

Consultants, Ltd. for Guyana Goldfields Inc., issued 2007.

GSEC, 2009: Environmental and Social Impact Assessment – Aurora Gold Mine, technical report



prepared by Ground Structures Engineering Consultants, Ltd. for Guyana Goldfields Inc.,

issued June 2009.

Guyana EPA, 2010: Environmental Permit 20090114-GGIOO, issued by Guyana Environmental

Protection Agency, September 28, 2010.

ICMC, 2009: International Cyanide Management Institute - Auditor Guidance for Use of theGold Mining Operations Verification Protocol, International Cyanide Management

Institute, October 2009.

NewFields, 2008: IFC Public Health Technical Assistance Program for Guyana Goldfields, Phase

1”, technical report prepared by NewFields for Guyana Goldfields Inc., issued 2008.

NewFields, 2008: IFC Public Health Technical Assistance Program for Guyana Goldfields, Phase

2”, technical report prepared by NewFields for Guyana Goldfields Inc., issued 2009.

World Bank/IFC, 2007: Environmental, Health and Safety Guidelines for Mining, World

Bank/International Finance Corporation, issued December 10, 2007.

World Bank/IFC, 2012: IFC Performance Standards on Environmental and Social Sustainability,


AMEC October 2011, Meteorological Data Update, Aurora Feasibility Study, prepared for

Guyana Goldfields Inc.

AMEC October 2011, Phase 1 Groundwater Flow Model Update, Aurora Feasibility Study,prepared for Guyana Goldfields Inc.



AMEC, 2009: Updated Preliminary Economic Assessment (PEA), Aurora Gold Project, Guyana,

South America, NI 43-101 Technical Report, June 2009.

AMEC-Geomatrix June 2010, Earthquake Ground Motion Hazard Evaluation, Aurora Feasibility

Study, prepared for Guyana Goldfields Inc.

Barton, N., Lien, R. and Lunde, J., 1974: Engineering classification of rock masses for the design

of tunnel support. Rock Mechanics, 6 May, p. 186-236.

Bieniawski, Z.T., 1976: Rock mass classification in rock engineering. In exploration for rock

engineering, Proc.. of the Symp. on Exploration for Rock Engineering (ed. Z.T.

Bieniawski) A.A. Balkema, Rotterdam, 1, 97-106. Cape Town.

Bienwiaski, Z.T. ,1989: Engineering rock mass classifications. New York, Wiley. 251 pp.

Byers, A.R. and Dahlstrom, C.D.A., 1954: Geology and mineral deposits of the Amisk – Wildcat

Lakes area, 63L-9, 63L-16, Saskatchewan; Saskatchewan Energy and Mines, Geological

Report No. 14, pp. 140-142.

C di D A i ti 2007 D S f t G id li


ICOLD , 2009: Tropical Residual Soils as Dam Foundation and Fill Material. ICOLD Committee on

Materials for Fills Dams. January 2009. pp. 99.

Klohn Crippen Berger 2011, Aurora project – Preliminary ARD/ML Interpretation, prepared forGuyana Goldfields Inc.



Martin, C.D., Kaiser,P.K., and McCreath, D.R., 1999: Hoek-Brown parameters for predicting

depth of brittle failure around tunnels. Can. Geotech. J., 36(1):136-151.

MWH 2008, Pre-feasibility Study of the Cuyuni hydroelectric Project, prepared for Guyana

Goldfields Inc.

Nickson, S.N., 1991: Cable support guidelines for underground hard rock mine operations.

MASc. thesis, Dept. Min. & Min. Pro., University of British Columbia.

Potvin, Y., 1988: Empirical open stope design in Canada. Ph.D. thesis, Dept. Min. & Min. Pro.,

University of British Columbia.

Snowden, 2008: Guyana Goldfields: Aurora Project Preliminary Assessment. Project Number

7178. Report to Guyana Goldfields Ltd., NI43-101 Technical Report.

Snowden, 2008: Technical Assessment: Geotechnical Notes. Internal Memo by Snowden to

Guyana Goldfields Ltd., February 2008.

SRK, 2012: Compilation of Feasibility documents for the Aurora Gold Deposit, Guyana Goldfields

Ltd (i )


2 8 . 0 C E R T I F I C A T E O F Q U A L I F I E D P E R S O N S

This Technical Report was prepared by the following qualified persons, certificates and consents of



p p p y g q p ,

which are contained herein:

Name Title, Company Responsible for Items

Glen Cole, P.Geo Principal Resource Geologist

SRK Consulting (Canada) Inc.

6, 7, 8, 9, 10, 11, 12, 14

Jarek Jakubec, C.Eng. Principal ConsultantSRK Consulting (Canada) Inc.

15, 16, portions of 24

John Lambert, P.Geo. Principal Consultant

ENVIRON EC (Canada) Inc.

20.1 to 20.4 and co-author for

20.5 and 20.6

D. Erik Spiller Principal Metallurgist

Tetra Tech, Inc.

13, 17, 19, 21, and 22

Richard Tocher, P.E. Principal Civil Engineer

Tetra Tech, Inc.

1, 2, 3, 4, 5, 18, portions of 6,

20, 24, 25 and 26









CERTIFICATE AND CONSENT

John Lambert, P.Geo.

P R I N C I P A L CON S U L T A N T

ENV IRON EC (C ANA DA ) I N C



ENV IRON EC ( C ANA DA ) I N C .

100 Park Royal, Suite 200

West Vancouver, British Columbia, Canada

Email: [email protected]

To accompany the report entitled: “43‐101 Technical Report, Updated Feasibility Study, Aurora

Gold Project, Guyana, S.A.”, dated January 29th, 2013.

I, John Lambert, P.Geo., of North Vancouver, BC, do hereby certify:

1) I am a Principal Consultant with ENVIRON EC (Canada) Inc. with a business address at Suite

200, 100 Park Royal West Vancouver, British Columbia, Canada;

2) I am a graduate of University of Hull, UK with a B.SC in Geology in 1973 and University of

Durham, UK with a M.Sc in Engineering Geology in 1974, and have practiced my profession

continuously since 1974. My work has involved engineering geology, environmental site

assessment, contaminated site investigation/remediation, environmental auditing and

environmental management systems for 37 years and consulting on cyanide code

TETRA TECH

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APPENDIX A

Letter from Guyana Goldfields Inc. Lawyer, March 4, 2012





APPENDIX B



APPENDIX B

Analytical Quality Control Data and Relative Precision Charts

(for May 2011 to May 2012 data)

Prepared bySKR Consulting



Time series plots for Field Blanks and Certified Reference Materials assayed by Acme Analytical

Laboratory between May 2011 and May 2012

Project Aurora

Data Series May 31 2011-May 31 2012Data Type DDH Samples - Blanks & Standards

Commodity Au in ppm

Laboratory Acme Labs

CDN- CDN- CDN- CDN- CDN-

Statistics GS-2J GS-3G GS-5F GS-10C GS-11A

Sample Count 17 197 395 2 61

Expected Value 2.36 2.59 5.3 9.71 11.21Standard Deviation 0.2 0.18 0.36 0.65 0.87

Mean 2.4 2.72 5.34 9.84 11.06

Outside 2StdDev 0.0% 1.5% 0.5% 0.0% 1.6%



Laboratory Acme Labs

Analytical Method Fire Assay - AAS and GRAV finish

Detection Limit 0.005 ppm

Outside 2StdDev 0.0% 1.5% 0.5% 0.0% 1.6%

Below 2StdDev 0 1 2 0 1

Above 2StdDev 0 2 0 0 0

1.7

1.9

2.1

2.3

2.5

2.7

2.9

3.1

2/25/2012 2/28/2012 3/9/2012 3/15/2012 3/27/2012 4/3/2012 4/25/2012 4/25/2012 5/31/2012

G o l d A s s a y ( p p m )

Samples (Time Series)

Time Series for CDN-GS-2J(Acme Labs; May 31 2011-May 31 2012 DDH Samples)

CDN-GS-2J

Expected Value

+2StdDev

-2StdDev

N = 17

1.8

2

2.2

2.4

2.6

2.8

3

3.2

3.4

5/31/2011 7/11/2011 7/20/2011 7/25/2011 8/2/2011 8/23/2011 10/2/2011 11/2/2011

G o l d A s s a y ( p p m )

Samples (Time Series)

Time Series for CDN-GS-3G(Acme Labs; May 31 2011-May 31 2012 DDH Samples)

CDN-GS-3G

Expected Value

+2StdDev

-2StdDev

N = 197

0.668 ppm Au.Likely referencematerial CDN-

GS-P7B.

6.5

7

Time Series for CDN-GS-5F(Acme Labs; May 31 2011-May 31 2012 DDH Samples)

CDN-GS-5F

Expected Value

+2StdDev

-2StdDev

N = 395

11

12

Time Series for CDN-GS-10C(Acme Labs; May 31 2011-May 31 2012 DDH Samples)

CDN-GS-10C

Expected Value

+2StdDev

-2StdDev

N = 2

Bias Charts, Quantile-Quantile and Relative Precision Plots for Field Duplicates assayed by Acme

Analytical Laboratory between May 2011 and May 2012

Project Aurora

Data Series May 31 2011 - May 31 2012

Data Type DDH Samples

Statistics Original Field Duplicate

Sample Count 988 988Minimum Value 0.00 0.00

Maximum Value 25.40 37.50

Mean 0 58 0 61



Data Type DDH Samples

Commodity Au in ppm

Analytical Method Fire Assay - AAS and GRAV finish

Detection Limit 0.005 ppm

Original Dataset Original Ass ays

Paired Dataset Field Duplicate Assays

Mean 0.58 0.61

Median 0.02 0.02

Standard Error 0.06 0.07

Standard Deviation 1.79 2.14

Correlation Coefficient 0.9037

Pairs ≤ 10% HARD 46.3%

y = 1.0759xR² = 0.8166

0

5

10

15

20

25

30

35

40

0 5 10 15 20 25 30 35 40

F i e l d D u p l i c a t e A s s a y s ( A u p p m )

Original Assays (Au ppm)

Bias Chart Field Duplicate Assay Pairs (0-40 ppm Au)

(Acme Labs; DDH Samples)

May 31 2011 -M ay 31 2012

+10%

-10%

N = 988 pairs

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

H A R D ( % )

Rank

Ranked Half Absolute Relative Deviation Plot

(Acme Labs; DDH Samples)

Au assayN = 988 pairs

46.3%

5

Bias Chart Field Duplicate Assay Pairs (0-5 ppm Au)(Acme Labs; DDH Samples)

N = 988 pairs100%

Mean versus Half Relative Deviation Plot(Acme Labs; DDH Samples)

Au assayN = 988 pairs



APPENDIX C

Modeled Variograms

Prepared bySRK Consulting

























APPENDIX D

Predictions of Groundwater Inflow

to Sublevel Retreat Mining

Prepared byItaska Denver, Inc.

PREDICTIONS OF GROUNDWATER INFLOW

TO SUBLEVEL RETREAT MINING

AT

GUYANA GOLDFIELDS AURORA MINE

GUYANA SOUTH AMERICA



GUYANA, SOUTH AMERICA

Prepared

for

SRK, Inc.

by

Itasca Denver, Inc.

143 Union Boulevard, Suite 525

Lakewood, Colorado

TABLE OF CONTENTS

Page

LIST OF FIGURES....................................................................................................................................... iv

LIST OF TABLES.......................................................................................................................................... v



LIST OF ABBREVIATIONS .......................................................................................................................... vi

EXECUTIVE SUMMARY ............................................................................................................................vii

1.0 INTRODUCTION ................................................................................................................................ 1

2.0 BACKGROUND .................................................................................................................................. 2

2.1 SITE CONDITIONS ..................................................................................................................... 2

2.2 PREVIOUS HYDROGEOLOGIC INVESTIGATIONS ...................................................................... 3

2.2.1 Summary of Field Programs ....................................................................................... 3

2.2.2 Analysis and Summary of Data from Previous Investigations ................................... 5

3.0 CONCEPTUAL HYDROGEOLOGIC MODEL ........................................................................................ 9

3.1 GENERAL HYDROGEOLOGIC SETTING ..................................................................................... 9

3.1.1 Climate ........................................................................................................................ 9

3.1.2 Topography, Drainage, and Vegetation ..................................................................... 9

3.1.3 Hydrologic Study Area .............................................................................................. 10

3.2 DESCRIPTION OF HYDROSTRATIGRAPHIC UNITS .................................................................. 10

TABLE OF CONTENTS

(continued)

Page

4.6.3 Simulation of Underground Sublevel Retreat Mining ............................................. 19

4.6.4 Simulation of Zone of Relaxation ............................................................................. 19



5.0 PREDICTIVE SIMULATIONS ............................................................................................................. 20

5.1 DISCUSSION OF KEY SIMPLIFICATIONS/ASSUMPTIONS IN THE MODEL PREDICTION ......... 20

5.2 SUMMARY OF MODEL SCENARIOS ....................................................................................... 20

5.3 SUMMARY OF PREDICTED INFLOW RATES ........................................................................... 21

5.3.1 Base Case Scenario ................................................................................................... 21

5.3.2 Sensitivity Analysis .................................................................................................... 21

5.4 EXPORT OF PORE PRESSURES FROM GROUNDWATER FLOW MODEL ................................ 22

5.5 ESTIMATED RAINFALL TO UNDERGROUND WORKINGS ....................................................... 23

6.0 CONCLUSIONS AND RECOMMENDATIONS ................................................................................... 24

6.1 CONCLUSIONS ........................................................................................................................ 24 6.2 RECOMMENDATIONS ............................................................................................................ 24

7.0 REFERENCES ................................................................................................................................... 26

LIST OF FIGURES

1 Base Map for Guyana Site

2

Conceptual Hydrogeologic Model

3 Measured Drawdown during Pumping Test at TW-7

4 Measured K vs Depth from 2008 2012 Field Programs



4 Measured K h vs. Depth from 2008 -2012 Field Programs

5 Measured K h and Modeled K x in Alluvium and Saprolite Based on Distance from River

6 Measured K h and Modeled K x in Weathered Bedrock Based on Distance from River

7

Measured K h and Modeled K x Values in Bedrock8 Measured Groundwater Levels Over Time During Pumping Test

9 Calculated Drawdown at End of Pumping Test for Different Thicknesses of Assumed Permeable

Zones and K h Values

10 Plan View of Simulated Hydrogeologic Zone of Unconsolidated Deposit

11 Cross Section A-A’ Showing Vertical Discretization of Model and Simulated Hydrogeologic Units

12

Plan View of Simulated Hydrogeologic Zone of Weathered Bedrock

13 Plan View Showing Boundary Conditions and Recharge Zones

14

Simulated Water Table Under Steady-State Conditions

LIST OF TABLES

1 Water Levels Used for Steady-State Conditions

2

Summary of Maximum Drawdown After 53.5-hour Pumping Test

3 Hydraulic Parameters of Geologic Units in Groundwater Flow Model

4 Open Pits Excavation Schedule



4 Open Pits Excavation Schedule

5 Estimated Rainfall over Pita Area that Reports to Underground Workings

6 Estimated Volume of Water over Pit Area that Reports to Underground Workings During a

Storm Event with 25 Years Return Period

LIST OF ABBREVIATIONS

3-D three dimensional

HSA hydrologic study area

K hydraulic conductivity

K horizontal hydraulic conductivity



K h horizontal hydraulic conductivity

K x, K v, K z hydraulic conductivity value along x, y, and z directions

L/s liters per second

LOM life of the minem meters

m3

cubic meters

m/s meters per second

m/day meters per day

m3/ day cubic meters per day

mamsl meters above mean sea level

mbgs meters below ground surface

SLR sublevel retreat

EXECUTIVE SUMMARY

(Page 1 of 2)

1) Itasca developed a three-dimensional (3D) finite element flow model of the Aurora Mine

located 170 km west of Georgetown in Guyana, South America to predict groundwater

inflows into the proposed Aurora pits and underground workings, and pore pressure



inflows into the proposed Aurora pits and underground workings, and pore pressure

distributions associated with open pit and underground mining.

2) Itasca first compiled and analyzed the hydrogeologic data from previous investigations

and incorporated these data into a conceptual hydrogeologic model of the mine area.

Then, based on this conceptual hydrogeologic model, a finite-element groundwater flowmodel of the Aurora Mine area was constructed. The model simulated the Cuyuni River

within the model domain, the Aurora open pits, and the underground mine workings.

3) The hydrogeologic units incorporated into the model include the unconsolidated deposits,

weathered bedrock, and fresh bedrock. The hydrologic properties of these geologic units

were simulated based on the analysis of the packer test and pumping test data from field

investigations provided by AMEC and SRK and refined through model calibration.

4)

Based on the pumping test data analysis, Itasca hypothesizes that there likely exists a thinpermeable unit within the weathered bedrock.

5) The shear zones were simulated in the model’s sensitivity analysis.

6) A pre-mining steady-state simulation of the groundwater flow model was calibrated to

EXECUTIVE SUMMARY

(Page 2 of 2)

•

For the weathered bedrock, the measured K h values of the areas closer to theCuyuni River are generally greater than the K h values of the areas farther away from

the Cuyuni River.



• The pumping test indicates that there is likely a thin permeable zone that lies

between the Cuyuni River and the mining areas. This permeable zone is likely

associated with the weathered bedrock.

• Under the base case scenario, the predicted maximum inflow rate to Rory’s Knoll

pit is about 600 m3/day, to Aleck Hill is about 1,500 m

3/day, and to SLR workings is

about 2,000 m3/day.

• The predicted inflow is moderately sensitive to the horizontal anisotropy ratio of

hydraulic conductivity. Assuming the horizontal anisotropy ratio of 1.0 will increase

the predicted inflow rate to SLR workings from 2,000 m3/day to about 4,000

m3/day.

•

The predicted inflow is highly sensitive to the permeable nature of the shear zones.

• Under the short term intensive storm event of 25 years return period, the volumes

of water that report to SLR underground workings range from 20,000 to 32,000 m3.

12) Based on Itasca’s findings from this phase of the 3-D groundwater flow model Itasca

1.0

INTRODUCTION

This report describes the groundwater flow model constructed by Itasca Denver, Inc., (Itasca)for Guyana Goldfields’ planned Aurora Mine in Guyana, South America. The model was

constructed with the geologic and hydrogeologic data collected from the mine area from past



g g y g g p

investigations. The purposes of this model were to:

1) simulate the open pit and sublevel retreat (SLR) mining,

2) predict potential flow rates into open pits and the SLR mining, and

3) provide predicted pore pressure distribution as input to geomechnic model.

Preliminary results of the groundwater flow model were provided to SRK Consulting (Canada)

Inc., (SRK) in a technical memorandum dated 24 October 2012 (Itasca 2012). That

memorandum was prepared without simulation of zone of relaxation (discussed in Section 3.4)

and only presented the predicted inflow rates without detailed discussion of the data and

groundwater flow model as described in this report.

2.0

BACKGROUND

2.1

SITE CONDITIONS

The Aurora mining project is located 170 km west of Georgetown in Guyana, South America.



The project site is located adjacent to the Cuyuni River, one of Guyana’s major rivers. Due to

the proximity of the river to the proposed mine, three segments of berm structures (or man-

made dykes) were proposed by Tetra Tech (email communication) to minimize the effects of

potential flooding on the mining operations. Figure 1 shows the locations of the planned open

pit and underground workings, Cuyuni River, and the proposed man-made dykes. Also shown in

Figure 1 are the topography, borehole locations from past investigations, and the model

domain.

A typical cross section of the area, as illustrated in Figure 2, consists of the unconsolidated

deposits (alluvium followed by a residual soil and Saprolite) which overlays the weathered

bedrock (also known as Saprock or transition rock) and finally the fresh bedrock. The residual

il S li d S k d i d f h h i f b k hi h i l d

2.2

PREVIOUS HYDROGEOLOGIC INVESTIGATIONS

Packer tests or pumping tests were conducted during field investigations in 2009, 2010, 2011,and 2012. The field investigations from 2009, 2010, and 2011 were conducted or supervised by

AMEC (AMEC 2010; 2011; 2012) and the 2012 packer test was supervised by SRK.



( ; ; ) p p y

Figure 1 shows the locations of boreholes drilled in 2009 (with prefix = “BH09-“), 2010 (with

prefixes = “BH10-“ or “TW-“), 2011 (with prefix = “BH11-“), and 2012 (with prefix = “BH12-“).

Most of these holes are located within the mining area. The holes with measured water levels

are summarized in Table 1. Table 1 also summarizes top and bottom elevations of the well

screens or the elevations at which the vibrating wire transducers (VWT) were installed. For the

open boreholes, the top and bottom elevations of the well screens refer to the top and bottom

of the opening section.

2.2.1 Summary of Field Programs

2009 Program

2010 Program

The 2010 program was performed by AMEC (AMEC 2011a) to provide horizontal hydraulic

conductivity (K h) values in the area between the proposed pit and the Cuyuni River. The

hydraulic testing in 2010 included both packer and pumping tests. Packer tests were conducted

i th f ll i l ti



in the following locations:

• Along the stretch of the proposed berm structure (man-made dykes): BH10-01 (TW-

1), BH10-02 (TW-2), BH10-03 (TW-3), BH10-04 (TW-4), BH10-05 (TW-5), BH10-06

(TW-6), TW-7 (Pumping Well), and TW-8

•

In the vicinity of the proposed open pit: BH10-RK-RMU-05 and BH10-RK-RMU-06

Short-term pumping tests and flow profiling were conducted to determine the inflow zones in

TW-2, TW-3, TW-4, TW-6, TW-7d, TW-8 and TW-8d. The short-term pumping tests were

conducted by pumping from one of the test wells and monitoring water levels in the nearbymonitoring wells for periods of up to 100 minutes (AMEC 2011a).

A longer 53.5-hour pumping test was also conducted by AMEC at TW-7 from December 10-12,

The shaft hole is located on the southeast side of Rory’s Knoll (Figure 1). The initial plan was to

conduct packer tests through the entire hole at 50 m intervals, but due to the limitation of the

testing equipment, packer tests were performed progressively at larger intervals below the

depth of 252 m below ground surface (mbgs) (AMEC 2012).



BH11-23 and BH11-24(C) are located between the proposed mine and the Cuyuni River (Figure

1). According to AMEC (2012), BH-11-24 is located within the shear zones initially identified by

SRK and BH11-23 is located between the shear zones and the Cuyuni River, north of the Rory’s

Knoll area (AMEC 2012).

2012 Program

The 2012 packer tests were conducted by SRK. The data from the packer tests were provided to

Itasca for hydraulic conductivity analysis. Except for BH12-SLC-02 and BH12-SLC-03, all other

holes are less than 50 m deep.

2.2.2

Analysis and Summary of Data from Previous Investigations

unconsolidated deposits with depth for different distance groups. Based on the available data,

the K h values were divided into three groups according to their distance from the bank of the

Cuyuni River:

• less than 200 m



• between 200 to 400 m

• between 400 and 1400 m

Figure 5 suggests that there is no clear spatial trend of K h in the unconsolidated deposits in

relation to the distance from the Cuyuni River.

Figure 6 shows the distribution of measured K h values in the weathered bedrock unit from the

bank of the Cuyuni River. Figure 6 shows that, except for one measured K h value in the 1400-m-

distance group and large ranges of measured K h values in the 2000-m-distance group, the

higher K h values are generally observed in the area that is closer to the Cuyuni River.

Figure 7a shows that, for the bedrock geologic units, there is no clear trend in the measured Kh

Therefore, there is no measured K h value from the shear zones. Nonetheless, the measured K h

values from both holes are generally low with a majority value of 10-4

m/day.

Observations from Pumping Test

In 2011 AMEC conducted several short-term pumping tests and one 53 5-hour pumping test



In 2011, AMEC conducted several short term pumping tests and one 53.5 hour pumping test

(AMEC 2011a). AMEC also conducted flow profiling in seven boreholes listed in Table 2. The

profiling results indicate that the water producing zone in most of the holes is between 15 to 60

mbgs. Based on AMEC's observation that "in test wells close to the river, the largest waterproducing zones were often located near the base of the casing, close to the bedrock contact

with the Saprolite", Itasca postulates that the contact zone, which was simulated in the model

as part of the weathered bedrock unit, could be more permeable than the other geologic units.

This is further discussed in the following paragraphs.

As shown in Table 2, pumping well TW-7 is an open hole, as well as the majority of the

monitoring wells. Subsequently, the pumping test was not a well-controlled test because it did

not target specific geologic units; however, in combination with the flow profiling, the pumping

m/day, the estimated drawdown at approximately 1000 m from the pumping well would be

smaller than the measured drawdown.

Figure 9b shows the calculated drawdown with distance from the pumping well assuming the

permeable unit is 20 m thick. As shown in Figure 9b, using the Theis formula, the estimated



drawdown at 1000 m from the pumping well would be smaller than the measured drawdown

for all ranges of the K h values.

Though the combination of thicknesses of the permeable unit and the assumed K h value is not

unique, Figure 9 demonstrates that, in order for the drawdown to propagate over 1000 m from

the pumping well, the permeable unit is likely to be relatively thin with high K h values (in the

range of 50 m/day). It is unlikely that all of these monitoring wells are connected by a

permeable feature, such as regional fault or shear zone, because (1) there is no shear zone

identified at the location of the pumping well, and (2) the flow profiling consistently shows that

the water-producing zones occur at the contact between the weathered and fresh bedrock,

which was penetrated by all the open boreholes.

3.0

CONCEPTUAL HYDROGEOLOGIC MODEL

The geologic setting (Figure 2) was briefly discussed in Section 2.1. Figure 2 illustrates theconceptual hydrogeologic model of the site along the north-south section. The major

hydrogeologic components include recharge, rivers, distribution of different geologic units, the



presence of shear zones, and hydraulic stresses induced by open-pit and underground mining.

The conceptual hydrogeologic model was developed based on the analysis of the site

conditions and data observed during the field investigation.

3.1

GENERAL HYDROGEOLOGIC SETTING

3.1.1

Climate

Aurora is located in a hot, humid, tropical environment with a high annual rainfall of

approximately 2450 mm. The annual evaporation is 1342 mm (AMEC 2011b). The typical rainy

seasons are from mid-April to mid-August and from mid-November to the end of January. There

is no data available on the local aquifer recharge rate; however, the recharge to the low-

3.1.3

Hydrologic Study Area

Figure 1 shows the Hydrologic Study Area (HSA) defined for the groundwater model of the

Aurora mine. The model boundaries were selected in such a way that the hydraulic stresses

induced by mining operations will not propagate to the model boundary. The Cuyuni River is



also located in the HSA. Several shear zones were identified within the HSA.

3.2

DESCRIPTION OF HYDROSTRATIGRAPHIC UNITS

The general stratigraphy of the site, from ground surface to below, consists of the following

major units:

1) Unconsolidated Deposits: This unit consists of alluvium and Saprolite (rock

weathered to a soil, but retaining the original structure of the parent rock)

2)

Weathered Rock: This unit consists mostly of the Saprock geologic unit

3) Fresh Bedrock: This unit consists of granite and meta-volcanic rock

The alluvium is generally thin (approximately 2 to 4 m). In some local granite highland areas and

Fresh bedrock at the site comprises granite, meta-sedimentary and meta-volcanic rocks of the

Cuyuni Formation. These different bedrock units were assumed to have similar hydraulic

conductivity values and are considered as one geologic unit in the model.

3.3

GEOLOGIC STRUCTURES



Shear zones were identified at the site. Some of the shear zones will be encountered by open-

pit and underground mining. As discussed in Section 2.2.2, no K h has been measured in the

shear zones.

3.4

MINING ACTIVITIES

The Aurora open pits, other than Rory’s Knoll pit, and SLR mining will proceed simultaneously.

SLR mining will only start after the completion of Rory’s Knoll pit. The SLR mining will cause the

disturbance to the rock surrounding the Rory’s Knoll pit wall and SLR mine workings. This

disturbance zone is defined in the groundwater flow model as the zone of relaxation (ZOR) and

simulated as a more permeable zone than the in-situ rock. Because the inflow to open pits and

4.0

GROUNDWATER FLOW MODEL

The groundwater flow model constructed for this investigation utilizes the commercial,numerical code MINEDW ™ (Azrag et al. 1998) developed by Itasca, which solves 3-D

groundwater flow problems with an unconfined (or phreatic) surface using the finite-element



method. The modeling code has been verified by Sandia National Laboratory (1998) and is used

at numerous mining hydrogeologic projects throughout the world.

4.1 MODEL DOMAIN AND DISCRETIZATION

The finite-element grid in plan view is shown in Figure 10. The finite-element discretization is

the finest (20 m) in the pit area to represent the geometry of the pit and the hydrogeology in

detail. The mesh gradually increases in size toward the boundaries of the model. For the base

case scenario, the shear zones were simulated with the same hydrogeologic parameters as thein-situ rock. The impacts of these shear zones on the inflow rate was simulated in the sensitivity

analysis. Therefore, the shear zones were represented as bands of finely discretized finite

elements as shown in Figure 10

4.2

SIMULATION OF HYDROGEOLOGIC SETTINGS

4.2.1 Simulation of Hydrogeologic Units

The hydrostratigraphic units simulated in the groundwater flow model were illustrated in the

conceptual hydrogeologic model (Figure 2). The geologic models provided by Guyana and SRK



p y g g ( g ) g g p y y

were used as the basis for the model construction, specifically, the model-defined top elevation

of the fresh bedrock. By assuming that the weathered bedrock has a uniform thickness of 5 m,

Itasca then defined the thickness of the alluvium/Saprolite unit from the top of the weatheredbedrock and ground surface.

The simulation of the hydrogeologic units is briefly summarized below:

Unconsolidated Deposits: This unit is a combination of alluvium and Saprolite. It is

simulated with two model layers. As shown in Figure 10, this unit was laterally simulatedwith two zones of different hydraulic conductivity values (hereafter referred to as

“hydrogeologic zones”) based on the distance from the Cuyuni River bank. This

simulation of two different hydrogeologic zones was mainly derived from the steady-

state model calibration and from the limited measured K h values (as shown in Figure 5).

4.2.2

Simulation of Man-Made Dykes

The representation of the dykes in the model is shown in Figures 10 to 12. These dykes are

assumed to extend from the ground surface to the top of the fresh bedrock. In the model, the

dykes were assumed to be constructed with low-permeability material.



4.3

MODEL BOUNDARIES

4.3.1

Recharge from Precipitation

Two recharge zones with slightly different recharge values were simulated according to the

ground surface elevation as shown in Figure 13. Based on the steady-state model calibration,

the recharge value was assumed to be 0.0206 and 0.0263 mm/day, respectively for the model

area that is lower and higher than 90 mamsl. In the predictive simulation, the recharge was

assumed to be constant over the entire model simulation. Due to its low-permeability nature,

no recharge was applied to the man-made dykes.

4 3 2 Variable Flux Boundary Condition

4.4

SIMULATION OF PRE-MINING CONDITIONS/STEADY-STATE CALIBRATION

Groundwater flow model simulations under steady-state conditions were conducted to

establish baseline groundwater levels. The simulated groundwater levels are compared with

the measured groundwater levels from various monitoring boreholes whose locations are

h i Fi 14 Fi 14 h th t d t fl f b th th th d th



shown in Figure 14. Figure 14 shows that groundwater flows from both the north and south

toward the river.

A comparison between simulated and the limited measured water levels shows that the model

generally matches the measured groundwater levels. The “quality line” in Figure 15 shows one

method to compare between the measured and simulated groundwater levels. As shown in

Figure 15, the trend in the measured groundwater levels generally agrees with that of the

simulated values.

4.5

SIMULATION OF PUMPING TEST

As discussed in Section 2.2.2, the 53.5-hour pumping test conducted by AMEC in December

approximately 1000 m from the pumping well. The simulated drawdown contours in the

weathered bedrock model layer are shown in Figure 16. In general, the simulated drawdown

follows the trend observed at the site. By assuming the thickness of the weathered bedrock is 5m, however, the model could not produce a 0.2 to 0.3 m drawdown at approximately 1000 m

from the pumping well as observed during the pumping test. No further attempt was made to



reduce the layer thickness of the permeable unit for the following two reasons:

1) the objective of the model simulation was not to calibrate the model to the pumping

test but rather to provide further understanding of the groundwater conditions atthe site. This objective was judged to have been achieved; and

2) the analytical solution has already demonstrated that, by decreasing the thickness of

the permeable zone, the estimated drawdown could reach about 0.3 m at

approximately 1000 m from the pumping well. By assuming a thicker permeable

zone than it could potentially be, the model would predict slightly more conservative

inflow rates.

Based on the model simulations of the pumping test, Itasca observed the following:

1) There could be a thin permeable unit between the pit and the Cuyuni River. In the

4.6

SIMULATION OF MINING

4.6.1 Open Pit Mine

Figure 18 shows the open pits and underground mine development. Five open pits were

simulated in the current model. Because there is no yearly pit plan, Itasca assumed that the



open pits will be excavated linearly according to yearly pit bottom elevation provided by SRK (G.

Carlson, email communication) which is summarized in Table 4.

Excavation of the open pit was simulated with drain nodes (or with a zero pore-pressure

condition) according to the assumed schedule and the final configuration of the open pit. The

purpose of simulating the pit excavation is to estimate the volume of groundwater that will

seep into the pit over the life of the mine (LOM). Therefore, the drain nodes that represent pit

excavation were turned ‘on’ according to the mining plan schedule.

Drain nodes were used to simulate the discharge of groundwater at the pit wall by the

relationship:

1

3 2

D

D D K f CL

⋅

where

K = hydraulic conductivity of the drain node material [m/day],

D1 , D2 , D3 = length related to the size of the individual element to which any particular drain is



associated [m], and

f = a factor that accounts for the effect of non-Darcian flow, the actual size of the

excavation relative to the grid size, and the shape of the excavation. This value is

generally calibrated to the measured inflow rate.

In MINEDW , the CL (leakance factor) can be calculated based on either the K value of the

geologic units using the above equation, calibrated based on the observed inflow rate, or

assigned with a large value. In predicting inflow to the Rory’s Knoll open pit, a large leakance

factor value was used, which essentially allows for groundwater from rocks to discharge freely

to each drain node without considering the effects of non-Darcian flow. This assumption may

lead to a slightly conservative prediction (i.e., an over estimate) of the seepage rate to the pit.

4 6 2 Si l i f R

4.6.3

Simulation of Underground Sublevel Retreat Mining

SLR mining begins in Year 5. Based on the SLR schedule provided by SRK and the extent of SLR,

Itasca interpolated the schedule for each SLR mine level in the model to represent progressive

mining over time. The top and bottom elevations of SLR mining are -70 and -970 mamsl,

respectively Underground mining was also simulated with drain nodes using the same



respectively. Underground mining was also simulated with drain nodes using the same

approaches described in Section 4.6.1.

4.6.4

Simulation of Zone of Relaxation

The extent of ZOR was provided by SRK based on its geomechanic model simulations. Because

the ZOR propagates both laterally and vertically as mining proceeds, SRK provides the ZOR

extent for every 50 m SLR mining stage (J. Severin, email communication). The development of

the ZOR over the LOM was simulated in the groundwater flow model by increasing the K value

of the rock within the ZOR. In the groundwater flow model, the K value of ZOR was assumed to

be 0.05 m/day (Table 3), which is one to two order(s) of magnitudes higher than the majority of

fresh bedrock.

5.0

PREDICTIVE SIMULATIONS

5.1

DISCUSSION OF KEY SIMPLIFICATIONS/ASSUMPTIONS IN THE MODEL PREDICTION

The following are key simplifications regarding the model predictions:



1) Shear zones were simulated to have similar hydrogeologic properties as the in-situ

bedrock. There are no measured data to demonstrate whether the shear zones are

permeable or not. A permeable shear zone could significantly increase the inflow

rates to both the open pit and the underground mine. Therefore, in the sensitivity

analysis, the effect of shear zones on the predicted inflow was evaluated.

2) Based on the analysis of the pumping test, Itasca hypothesized that a thin

permeable zone exists along the river band. The extent of this permeable zone is

unknown. Based on the model simulations of the 53.5-hour pumping test, this thin,

permeable zone was assumed to exist within 200 m along the river. This hypothesis

requires further confirmation from future field investigations.

3) The effect of surface water runoff and direct precipitation from rain/storm events on

the open pits and underground pumping requirements are not reflected in the

predicted inflow rates. Direct precipitation over the foot print of the Rory’s Knoll

open pit and surface runoff to the pit will both report to the underground workings

Scenario 4: This scenario assumed that the K x value of the shear zones is 0.5 m/day,

which is 10 times greater than the K x value of the upper bedrock in

Scenario 1.

5.3

SUMMARY OF PREDICTED INFLOW RATES

5 3 1 Base Case Scenario



5.3.1

Base Case Scenario

The predicted inflow rates to various open pits shown in Figure 19 leads to the following

observations:

1) After Year 6, the inflow to Rory’s Knoll pit will report to SLR mine workings.

2)

Among all open pits, the Aleck Hill Pit would encounter the highest inflow rate with

the maximum inflow rate of about 1,600 m3/day because of its large footprint and

depth.

3)

The predicted inflow rates to other open pits are generally less than 200 m3

/day.

Figure 20 shows the estimated inflow rate to the SLR workings over time. After the initial

increase due to the ramp development the predicted inflow rate increases from 1 700 m3/day

value of the shear zones. By arbitrarily assigning the K x value of the shear zones as 0.5 m/day,

the maximum inflow rates increase by almost 10 times in comparison to the base case scenario.

5.4

EXPORT OF PORE PRESSURES FROM GROUNDWATER FLOW MODEL

The pore pressure was required by SRK for the domain and intervals as shown in the following



The pore pressure was required by SRK for the domain and intervals as shown in the following

table. The minimum and maximum coordinates refer to the lower and upper corner of the

domain. The interval is the dimension of the blocks. Based on the block size and domain, the

number of block segments was derived.

Coordinates Minimum (m) Maximum (m) Interval (m)Number of Block

Segments

x 195143 197339.08 7.76 283

y 750352 752144.56 7.76 231

z -1800 108.96 7.76 246

The interpolated pore pressure for the above domain has a total of 16,274,336 records (or lines)

(284 × 232 × 247). Each line has the format:

The interpolated pore pressures from the groundwater flow model were provided to SRK for pre-

mining, end of open pit, and 17 SLR stages. Each SLR stage covers the depth interval of 50 m.

5.5

ESTIMATED RAINFALL TO UNDERGROUND WORKINGS

The rainfall that reports to the underground workings is estimated based on the rainfall data



p g g

provided by AMEC (2011b). Table 5 summarizes the potential rainfall that reports to the

underground working area. The calculations were based on the average monthly precipitation and

evaporation rate from the Timehri Climate Station (AMEC 2011b). As shown in the table, the

rainfall to the underground workings contains both runoff from Rory’s Knoll pit wall and direct

precipitation to the underground workings. Table 5 did not include the runoff from the watershed

catchment by assuming that engineering measures will be taken to avoid runoff from the

catchment area.

The total rainfall was calculated based on the following assumptions:

1) Evaporation was assumed to occur on the pit wall.

2) B d I ’ j i 70% f h ff f h i ll d

6.0

CONCLUSIONS AND RECOMMENDATIONS

6.1

CONCLUSIONS

Based on the data analysis and model simulations, Itasca concludes the following:



1) The measured K h values of the fresh bedrock decreases with depth.

2) For the weathered bedrock, the measured K h values in the areas closer to the Cuyuni

River are generally greater than those in the areas farther away from the CuyuniRiver.

3) The pumping test indicates that there is likely a thin permeable zone that lies

between the Cuyuni River and mining areas. This permeable zone is likely associated

with the weathered bedrock.

4) Under the base case scenario, the predicted maximum inflow rate to Rory’s Knoll pit

is about 600 m3/day, to Aleck Hill is about 1,500 m

3/day, and to SLR workings is

about 2,000 m3/day.

5) The predicted inflow is moderately sensitive to the horizontal anisotropy ratio of

hydraulic conductivity. Assuming the horizontal anisotropy ratio of 1.0 will increase

the predicted inflow rate to SLR workings from 2 000 m3/day to about 4 000 m

3/day

2) For the same reason as above, the second pumping test should be conducted in the

weathered bedrock unit.

3) The third pumping test should be conducted in the shear zones. The well should be

screened from the top of the shear zone (assumed to be the top of the bedrock) and

50 m below the top of the fresh bedrock. This test would provide data to evaluate

the permeable nature of the shear zones.



Prior to these pumping tests, multi-level piezometers should be located and installed at five to

six locations. The changes in water levels from these multi-level piezometers are valuable data

in analyzing the groundwater conditions at the site. The locations of pumping wells and multi-

level piezometers will be jointly decided by Guyana, SRK, and Itasca.

After the completion of the pumping tests, the groundwater flow model should be recalibrated

to these pumping tests and, subsequently, used to simulate inflow rates to the mines.

7.0

REFERENCES

AMEC. 2010. Geotechnical Investigation for the feasibility study Aurora Gold Project, Guyana:

Report submitted to Guyana Goldfields Inc. August 2010.

AMEC. 2011a. Pumping test results from the 2010 Testing program for Aurora Gold Project:

Memorandum submitted to Guyana Goldfields Inc April 2011



Memorandum submitted to Guyana Goldfields Inc. April 2011.

AMEC. 2011b. Meteorological data update, Aurora feasibility study, Guyana Goldfields, Guyana.

Draft memorandum submitted to Guyana Goldfields Inc. 4 October2011.

AMEC. 2012. Review of Phase 2 deep packer testing 2011 Aurora feasibility study, Guyana

Goldfields, Guyana: Memorandum submitted to Guyana Goldfields Inc. February 2012.

Itasca. 2012. Predicted Inflow Rate to Sublevel Retreat Mining at Aurora Mine. Technical

memorandum submitted to SRK Consulting (Canada) Inc., by Itasca Denver, Inc. 24 October.











































TABLE 1

Water Levels Used for Steady-State Calibration

Borehole IDEasting

(m)

Northing

(m)

Ground

Elevation

(mamsl)

Bottom of

Screen

(mamsl)

Top of

Screen

(mamsl)

Screen

Interval

(m)

Water

Level

(mamsl)

Instrument Type Geologic Unit

BH09-01 195742.5 751901.3 55.3 47.0 53.0 VW Ash Tuff and Tuff

BH09-02 196089.9 751909.4 54.9 45.7 49.7 VW Tuff/volvanic sediments

BH09-03A 196408.8 751903.4 56.0 31.9 33.4 1.5 53.2 Stand pipe Saprolite w/rock

BH09-03B 196407.4 751903.2 55.9 46.8 48.3 1.5 53.7 Stand pipe Alluvial w/Saprolite

BH09-04A 196717.0 751929.1 56.5 52.3 53.8 1.5 53.3 Stand pipe Alluvial

BH09-04B 196713.1 751932.0 56.4 38.1 44.2 6.1 52.2 Stand pipe Saprolite Weathered Bedrock

BH09-04C 196714.2 751928.3 56.6 49.0 53.8 VW NA(1)

BH09-05 197006.0 751887.7 56.8 29.5 49.7 VW Volcanic sediments

BH09-08A 197019.3 751714.3 53.5 42.7 44.2 1.5 53.2 Stand pipe Saprolite

BH09-08B 197016.9 751716.0 53.4 49.2 VW Fresh Bedrock

BH09-11 195583.5 750892.2 63.4 -13.4 60.9 VW Fresh Bedrock

BH09-15 197834 0 751122 0 55 5 37 2 40 3 3 1 53 6 Stand pipe Saprolite



BH09 15 197834.0 751122.0 55.5 37.2 40.3 3.1 53.6 Stand pipe Saprolite

BH09-16 197958.9 751195.3 56.9 34.3 37.4 3.1 50.4 Stand pipe Saprolite

BH09-19 193313.4 746959.1 101.6 76.9 80.0 3.1 101.1 Stand pipe Weathered Bedrock

BH09-20 193284.5 747214.8 92.7 62.7 77.7 VW Fresh Bedrock

BH09-21A 193259.2 747457.5 81.5 60.4 63.4 3.0 71.7 Stand pipe Saprolite

BH09-22 193234.3 747643.6 75.1 70.1 73.1 3.0 72.9 Stand pipe Saprolite w/rockBH09-23 193209.4 747846.3 118.2 94.4 97.5 3.1 95.5 Stand pipe NA

(1)

BH09-24 193135.8 748110.9 87.5 62.7 78.4 VW Fresh Bedrock

BH09-25 195794.7 747754.8 86.6 69.6 72.6 3.0 71.8 Stand pipe Saprolite w/rock



BH09-30A 194630.3 751205.3 61.0 46.1 55.6 VW NA(1)


BH09-32 194553.0 750767.3 58.2 43.4 57.6 VW Saprolite

BH10-1 (TW 1) 195740.1 751906.0 50.4 -149.6 38.4 188.0 49.1 Meta-volcanics

TW-8 195738.3 751904.3 50.4 -171.6 35.4 206.0 49.2 Fresh Bedrock

TW-8a 195722.4 751886.3 50.4 -200.6 35.4 236.0 49.6 Fresh BedrockTW-8b 195758.6 751886.4 49.7 -202.3 37.7 240.0 49.1 Fresh Bedrock

TW-8c 195722.3 751922.3 50.7 -197.3 35.7 233.0 49.6 Fresh Bedrock

TW-8d 195758.4 751922.4 50.6 -201.4 35.6 237.0 49.4 Fresh Bedrock

Note: 1) Geologic unit is not available.

Open Hole

TABLE 2

Summary of Maximum Drawdown After 53.5-hour Pumping Test

Borehole IDEasting

(m)

Northing

(m)

Distance from

Pumping Well

(m)

Measured

Drawdown

(m)

Top of Screen

(mamsl)

Bottom of Screen

(mamsl)

Instrument

TypeGeologic Unit

BH09-02 196089.9 751909.4 779.2 0.0 45.7 45.7 VW Tuff/volvanic sediments

BH09-03A 196408.8 751903.4 460.5 1.0 33.4 31.9 Stand pipe Saprolite w/rock

BH09-03B 196407.4 751903.2 461.9 0.6 48.3 46.8 Stand pipe Alluvial w/Saprolite

BH09-04B 196713.1 751932.0 167.2 1.8 44.2 38.1 Stand pipe Saprolite/weathered bedrock

BH09-04C 196714.2 751928.3 164.8 1.6 49.0 49.0 VW NA(1)

BH09-05 197006.0 751887.7 139.1 2.9 29.5 29.5 VW Volcanic sediments

BH09-11 195583.5 750892.2 1613.9 0.1 -13.4 -13.4 VW Fresh Bedrock

BH09-RK-RMP-01A 196732.4 751734.4 190.8 0.9 NA(2) NA(2) Stand pipe NA(1)

BH09-RK-RMU-01A 196685.0 751824.5 187.9 2.7 NA(2)

NA(2) Stand pipe NA

(1)

TW-7 196868.1 751869.3 0.0 3.9 36.0 -166.5

TW-7a 196852.9 751851.4 23.4 2.8 32.9 -200.1

TW-7b 196888.4 751851.4 27.1 2.7 32.6 -199.4

TW-7c 196852.4 751887.7 24.2 3.1 29.7 -199.3

TW-7d* 196888.5 751887.6 27.4 3.0 33.0 -200.0

TW 8* 195738 3 751904 3 1130 3 0 3 35 4 171 6



TW-8* 195738.3 751904.3 1130.3 0.3 35.4 -171.6

TW-8a 195722.4 751886.3 1145.8 0.3 35.4 -200.6

TW-8b 195758.6 751886.4 1109.7 37.7 -202.3

TW-8c 195722.3 751922.3 1147.0 0.3 35.7 -197.3

TW-8d* 195758.4 751922.4 1111.0 0.3 35.6 -201.4

TW-1 195740.4 751904.3 1128.2 38.4 -149.6TW-2* 196420.3 751728.3 469.2 0.1 12.4 -133.6

TW-3* 196555.8 751863.8 312.6 1.3 39.5 -157.5

TW-4* 196692.9 751786.7 193.6 0.9 26.2 -151.8

TW-5 196870.0 751869.3 2.0 3.9 30.1 -190.9

TW-6* 197009.7 751855.8 141.8 3.1 26.9 -151.1

Notes: * Boreholes with flow profiling.

1) Geologic unit is not available.

2) Screen information is not available.

Open Hole Fresh Bedrock

TABLE 3

Hydraulic Parameters of Geologic Units in Groundwater Flow Model

K x K y K z

< 1400 m from the riverbank 1.0E-01 1.0E-01 1.0E-02 1.0E-05 2.0E-01


< 200 m from the riverbank 5.0E+01 5.0E+00 5.0E+00 1.0E-06 1.0E-02




Upper Bedrock: above -15 mamsl 5.0E-02 5.0E-03 5.0E-03 1.0E-06 5.0E-03

Middl B d k b t 15 l d 195 l 3 2E 02 3 2E 03 3 2E 03 1 0E 06 5 0E 03

Unconsolidated Deposits

Weathered Bedrock

Hydraulic Conductivity

(m/day)Specific

Storage

(m-1

)

Specific

Yield

( )

Formation/Unit



Middle Bedrock: between -15 mamsl and -195 mamsl 3.2E-02 3.2E-03 3.2E-03 1.0E-06 5.0E-03

Lower Bedrock: between -195 mamsl and -245 mamsl 4.7E-03 4.7E-04 4.7E-04 1.0E-06 5.0E-03

Deep Bedrock: below -245 mamsl 3.6E-04 3.6E-05 3.6E-05 1.0E-06 5.0E-03

Zone of Relaxation for prediction only 5.0E-02 5.0E-02 5.0E-02 5.0E-06 5.0E-03

Dyke for prediction only 1.0E-05 1.0E-05 1.0E-05 1.0E-06 5.0E-03

Shear Zone for sensitivity analysis only 5.0E-01 5.0E-02 5.0E-02 1.0E-05 5.0E-03

Fresh Bedrock

TABLE 4

Open Pits Excavation Schedule

Start End

SAP 25 1 3

PB1 -35 4 5

PB2 -95 4 6

Pit Name

Rory's Knoll

Excavation Schedule

(Production Years)Bottom Elevation

(mamsl)



SAP 30 1 4

PB1 -35 4 7

PB2 -110 4 14SAP 45 2 3

PB1 0 8 13

SAP 40 4 4

PB1 -60 11 15

Walcott Hill PB1 35 14 15

Mad Kiss

Aleck Hill

Aleck Hill North

TABLE 5

Estimated Rainfall over Pit Area that Reports to Underground Workings

Jan 207 6.80 85.4 2.81 379 53 432

Feb 102 3.35 89.3 2.93 40 26 66Mar 134 4.40 105.1 3.45 90 34 124

Apr 172 5.65 112.1 3.68 187 44 231

May 317 10.41 118.3 3.89 620 81 701

June 337 11.07 138.8 4.56 618 86 704

July 286 9.40 98.6 3.24 584 73 658

Aug 219 7.19 86.7 2.85 412 56 469

Total

Rainfall

Rainfall to Underground Workings

(m3/day)

Evaporation

Monthly

(mm/month)

Daily

(mm/day)

Run-off1, 2, 3

from

Rory's Knoll Pit Wall over

Area of 135,580 m2

Direct Precipitation

to UG over

Area of 7,800 m2

Month

Precipitation

Monthly

(mm/month)

Daily

(mm/day)



g

Sept 132 4.34 141.4 4.65 0 34 34

Oct 137 4.50 144.4 4.74 0 35 35

Nov 174 5.72 96.8 3.18 241 45 285

Dec 233 7.65 124.9 4.10 337 60 397

Notes: 1) The area excludes the 7,800 m2 of direct precipitation.

2) Evaporat ion was substracted f rom the precipi ta tion .

3) Assumes that 70% of the runoff from the pit wall reports to the underground workings.

TABLE 6

Estimated Volume of Water over Pit Area that Reports to

Underground Workings During a Storm Event with 25 Years Return Period

Volume of Rainfall to Underground Workings

During the Storm Event (m3)

3 40.3 16,926

Duration

(Hours) Rainfall1, 2, 3

to Rory's Knoll Pit Perimeter Area

of 140,000 m2

Rainfall Intensity

(mm/hour)



3 40.3 16,926

6 25 21,000

12 15.5 26,040

Notes: 1) The area does not include any runoff catchment area.

2) Evaporation was not included in this short term event.

3) Assumes that all precipitation reports to the underground workings.

APPENDIX E



Underground Electrical PowerDistribution Design for the

Aurora Gold Project

Prepared byTkaczuk & Associates Inc.

Tkaczuk & Associates Inc.



Underground Electrical PowerDistribution Design for the

1.0 Executive Summary

This report provides a design and capital cost estimate for the under ground portion of Aurora Gold Pro- ject underground electrical power distribution system and the main ventilation fans on surface.




The design and cost for the primary power supply on surface was completed by others, and was basedon utilizing Diesel generation.

The design included the power required for the under ground operation and they provided aconnection point for the underground power cables in the main electrical substation.

This package only covers the design and cost for the !backbone" of the underground powerdistribution system. The total electrical load for under ground is based on the development, production,ore handling, dewatering and infrastructure requirements.

The major electrical infrastructure included in this work package includes:

Substation at the main surface intake and exhaust fans.• Pole line on surface to the main fresh air fans.• The installation of three 13.8 Kv power cables from the main substation.• The installation of under ground switch rooms for the development and production require-

ments.

2.0 Assumptions and Exclusions

2.1 Assumptions

The following assumptions were used in the design and capital costs for both options




The following assumptions were used in the design and capital costs for both options:• There is sufficient 13.8 KV capacity at the surface substation

2.2 ExclusionsThe following are excluded from the design and capital costs for both options:

• All electric motors.• All excavations and cable holes between levels.

Any electrical cost for major infrastructure and the connection to the electrical backbone,including:

• Maintenance facilities.• Refuge Stations.

• Main dewatering stations (infrastructure only)• Ore and waste handling.• Backfill.

3.0 Electrical System Overview

3.1 System Buildout

Electrical power is required for the underground development and ongoing production of the Rory"s Knoll




Electrical power is required for the underground development and ongoing production of the Rory"s Knoll.

To facilitate development activities, independence from the production power infrastructure is required.

Each ramp, lateral and raise development crew will be provided electrical distribution equipment whichwill allow them to proceed at their own pace. A typical crew setup will consist of the following.• Medium voltage (13.8 kV) junction box• Medium voltage (13.8 kV) portable power cable• Mine portable substation (1 Mva, 13.8kV/460V)• Portable rack of ground fault protected starters configured to meet the needs of specific crews

(Jumbo, fans, face pumps, raise bore, etc…)

Crews will tie into permanent 13.8 kV infrastructure and advance as per schedule. Utilizing 13.8 kV pro-

vides the capability to develop much further before needing to establish a shorter tie. This allows perma-nent infrastructure to be constructed, tested and commissioned well in advance of the next leg of devel-opment.

3.1 Main Feeders

3.2 Portable Substations and Equipment

The SLR mining method allows the electrical distribution system to be arranged into 9 major portablesubstations (1 Mva) each servicing 4 sublevels. Where additional development takes place such asraise boring and ramp development additional substations can be installed as required. Smaller (500Kva) substations are used on the lightly loaded sublevels.

Equipment has been selected to minimize the number of different pieces. This provides a standardequipment fleet which can be easily managed, maintained and re-deployed on demand. Sufficient quanti-ties of equipment will be available to minimize the need to purchase additional capital equipment as themine develops. During latter stages of mine life there will be a surplus of electrical equipment which maybe salvaged from the upper levels refurbished and redeployed to lower levels.

All equipment starters are modular rack mounted with plugs Each starter is ground fault protected im




All equipment starters are modular rack mounted with plugs. Each starter is ground fault protected im-proving overall system reliability and safety.

4.0 Power Demand and Consumption

4.1 Energy Demand

Energy demand is based on loads identified in Table 1 with associated de rating factors Factors have




Energy demand is based on loads identified in Table 1 with associated de-rating factors. Factors havebeen selected based on anticipated equipment duty cycles required to support the mining method as wellas development and production schedules.

Table 1: Underground Electrical Loads

Description Installe Power De-rating Factor Load

Hp kW % kW

Development Jumbo/s 294 220 100 220

Production Drill/s 320 239 100 239

Emulsion Loader/s 276 206 100 206

Bolter 276 206 100 206Shotcrete/Transmixer/s 276 206 100 206

Refuge Stations 30 22 50 11

U/G Shop 50 37 75 28

52.578 1.301 0.716 0.507 1.014 4.995 5.102 5.717

62.658 1.471 0.802 0.507 1.014 5.317 5.437 6.065

72.731 1.776 0.888 0.622 1.244 5.883 6.016 6.771

82.784 1.945 0.900 0.622 1.244 6.116 6.251 7.008

92.833 2.115 0.900 0.737 1.473 6.450 6.585 7.456

102.878 2.285 0.900 0.852 1.703 6.779 6.914 7.901

112.920 2.454 0.900 0.852 1.703 6.991 7.126 8.112

122.778 2.454 0.900 0.852 1.703 6.849 6.984 7.971

132.814 2.454 0.900 0.852 1.703 6.885 7.020 8.007

142.654 2.454 0.861 0.852 1.703 6.692 6.821 7.802

152 654 2 454 0 198 0 852 1 703 6 638 6 758 7 729




2.654 2.454 0.198 0.852 1.703 6.638 6.758 7.729

Figure 1: Electrical Energy Demand

Electrical Power Demand Profile

5 000

6.000

7.000

8.000

9.000

MW Min

5.0 Power Generation

5.1 Requirements

A previous study "Power Generation Survey for the Aurora Gold Project" prepared by Martin Menard




A previous study Power Generation Survey for the Aurora Gold Project prepared by Martin MenardConsultant Inc. evaluated the use of Wartsila 12V32, 5.2MW heavy fuel oil power plants as a means ofproviding onsite power.

Based on the continued use of this type of equipment a minimum of two generating units will be requiredto satisfy the peak underground load.

6.0 Project Costs

6.1 Capital Costs

Costs associated with the acquisition installation and commissioning of electrical equipment are summa-




Costs associated with the acquisition, installation and commissioning of electrical equipment are summarized in Table 3.

Table 3: Electrical Infrastructure Costs - Total

Scope of Electrical Work Material Cost (USD) Construction (Mhrs.)

Raise Development Crew Equipment $176,324 0

Development Crew Equipment $702,306 0

Substations & Medium Voltage Cable $8,623,901 15,083

Misc. Construction Hardware $198,491 3,767

Low Voltage Cable $831,723 5,973

Total $10,532,746 24,822

The electrical distribution system will be built-out over the course of mine development and production.Capital costs will be distributed approximately as shown in Table 4.

Table 4: Electrical Infrastructure Costs - Year

Period Material Cost

Years USD

-2 $758,400

-1 $934,724

1 $803,6022 $803,602

3 $803,602

4 $803,602

5 $803,602




6 $803,602

7 $803,602

8 $803,602

9 $803,602

10 $803,602

11 $803,602

12 $0

13 $0

14 $0

15 $0

Total $10,532,746

It is anticipated that 3 to 4 electricians will be required on a continuous basis to support the build-out ofthe electrical distribution system over the course of the mine's life

6.2 Manpower and Maintenance Costs

Costs associated with the operation of the electrical distribution system consist of three main compo-nents: energy, maintenance material and maintenance labour. Manpower resources include 14 electricaltradespersons which is sufficient for the size of distribution system anticipated.

Maintenance material requirements are highly variable, given location, skill levels, management and op-erating conditions. A modern well organized operation can expect routine maintenance material costs inthe range of 10 to 20% of capital. This translates to an annual cost of $65,000 to $130,000.

6.3 Manpower and Maintenance Costs

Cost of electrical energy is shown in Table 5 and Figure 2 and is based on $0.265 as determined in thet d "P G ti S f th A G ld P j t" N l ti i f l t i i l d d




study "Power Generation Survey for the Aurora Gold Project". No escalation in fuel cost is included.

Table 5: Electrical Energy Cost

ear Total Demand En rgy Cost

Nominal MW By Year Cumulative

-21.018 $2,363,978 $2,363,978

-11.232 $2,860,796 $5,224,774

12.390 $5,549,247 $10,774,021

23.016 $7,001,185 $17,775,207

3 4.291 $9,961,479 $27,736,6854

4.644 $10,780,672 $38,517,3575

5.102 $11,844,308 $50,361,6656

5.437 $12,622,482 $62,984,148

Yearly and Cumulative Electrical Energy Cost

$150,000,000

$200,000,000

$250,000,000

U S D )




$-

$50,000,000

$100,000,000

1 3 5 7 9 11 13 15 17Year

D o

l l a r s ( U

$/Year

Accum. $

7.0 Risks and Opportunities

7.1 Risks

Table 5 and Figure 2 "Electrical Energy Cost" indicate a potential significant project risk associated with




g gy p g p jthe cost of fuel.

A related risk that will manifest itself as higher fuel costs is uncontrolled electrical load growth that nor-mally translates to higher electrical energy demand.

7.2 Opportunities

The mining method selected provides an opportunity to significantly reduce electrical infrastructure capitalcost by establishing an effective maintenance system. Activities should focus on downsizing equipment inmined out levels, refurbishing equipment wherever cost effective and redeploying equipment in newlyopened production areas. Potential exists for the re-use of up to 30% of the primary infrastructure.

An effective energy management system that incorporates on line monitoring of demand and load controlwill directly translate into fuel savings. The system should also incorporate procedures that require a cost-benefit analysis for proposed load increases.

Appendix: Single Line Diagrams
















APPENDIX F



Aurora Gold ProjectFeasibility Ventilation and

Refrigeration Study

Prepared byBBE Consulting.



AURORA GOLD PROJECT

FEASIBILITY VENTILATION ANDREFRIGERATION STUDY

Sub-Level Retreat (SLR) mining method

December 2012

BBE Report 6912

Rev 2

Aurora Project Guyana, Rory’s Knoll, Feasibility Ventilation and Refrigeration Study, BBE report 6912 December 2012, Rev 2

EXECUTIVE SUMMARY

The ventilation and cooling requirements for mining down to -979 Level at Rory’s Knoll have beenestimated. VUMA-network simulations have been carried out to estimate mine heat loads.Ventilation requirements are strongly influenced by the expected diesel fleet.

The ultimate needs are as follows.

Total ventilation Total air flow specification [D/C and U/C flow] 512 kg/s

Downcast ventilation Intake ramp 212 kg/s Fresh Air Raise 300 kg/s



Fresh Air Raise 300 kg/s

Ai r coo lers Main downcast raise surface air cooler duty 11 MWR

Upcast ventilation 4.0 m diameter upcast ventilation raise 256 kg/s 4.0 m diameter upcast ventilation raise 256 kg/s

The capital estimate for the major equipment is as follows:

Surface Main Fan Station[s]CAPEX provision for two main surface fan stations will be $6.8M.

Secondary Ventilation EquipmentCAPEX provision for auxiliary fans, ducting, refuge stations, early warning system will be $8.5M.


CONTENTS

1

INTRODUCTION AND SCOPE OF WORK................................................................................1

2

DESIGN CRITERIA .................................................................................................................... 2

2.1

Production Profile and Mine Layout ...................................................................................2

2.2 Ore Extraction Method .......................................................................................................4

2.3 Main Airway Velocity .......................................................................................................... 4

2.4

Friction Factors................................................................................................................... 4

2.5

Leakage and Commitments [Controlled and Uncontrolled] ................................................4

2.6

Development Ventilation ....................................................................................................4

2.7

Diesel Equipment ............................................................................................................... 5

2.8

Surface Ambient Design Condition ....................................................................................5

2.9

Geothermal Properties .......................................................................................................5



2.10

Production Face Conditions ...............................................................................................6

2.11

Ramp Conditions ................................................................................................................ 6

2.12

Water .................................................................................................................................. 6 2.13

Refuge Chambers ..............................................................................................................7

2.14

Second Outlets ................................................................................................................... 7

2.15

Airborne Pollutants [General Air] ........................................................................................7

2.16

Noise .................................................................................................................................. 7

2.17

Economic Parameters ........................................................................................................7

2.18 Workshops / Tyre Bays / Bulk Liquid and Oil Stores ..........................................................7

2.19 Fire Suppression for Vehicles and Machinery ....................................................................7

3

HEAT LOAD AND VENTILATION NETWORK ANALYSIS ......................................................8

3.1 Critical Snap-Shots............................................................................................................. 8

3.2 Heat Load Components .....................................................................................................8

3.3

Results of VUMA Models ...................................................................................................9

3 3 1 M i d th f i i b f f i ti i i d Y 3 d 9




TABLES

Table 1

Preliminary Production Schedule .................................................................................... 3

Table 2

Main Airway Velocities .................................................................................................... 4

Table 3

Main Airway Friction Factors ........................................................................................... 4

Table 4 Leakage Factors ............................................................................................................. 4

Table 5 Development Ventilation Factors .................................................................................... 4

Table 6

Final Diesel Fleet ............................................................................................................ 5

Table 7

Rock Properties .............................................................................................................. 6

Table 8

Main Motor List for Surface Refrigeration System ........................................................ 28

Table 9

Absorption Chiller Capital Costs ................................................................................... 31

Table 10

Capital Cost Payback – 5% Yearly Escalation .............................................................. 32

Table 11

Capital Cost Payback – No Escalation ......................................................................... 33



Table 12

Power Requirements of Main Components and Operating Costs ................................ 36

Table 13

Ventilation Officials’ Instruments ................................................................................... 38

Table 14

Early Warning System .................................................................................................. 38 Table 15

Civils ............................................................................................................................. 39

Table 16

Refrigeration Machines ................................................................................................. 39

Table 17

Mechanical and Electrical ............................................................................................. 39

Table 18

Main Fans and Secondary Ventilation .......................................................................... 40

FIGURES

Figure 1

Mine Layout .................................................................................................................... 2 Figure 2

VUMA snapshot of -354 Level wet-bulb temperatures before refrigeration .................... 9

Figure 3

VUMA snap-shot of -604 Level to -704 Level wet-bulb temperatures........................... 10

Figure 4

VUMA snap-shot of wet-bulb temperatures for the ultimate scenario ........................... 11

Figure 5 Regulator on non-producing levels 15


1 INTRODUCTION AND SCOPE OF WORK

BBE Consulting carried out a ventilation/refrigeration study for the Aurora Project in 2011 [miningto -1 230 Level, open stoping with cemented backfill with a production rate of 4 000 tpd] and asubsequent study for SRK Consulting [Toronto] [mining to -720 Level, open stoping with cementedbackfill with a production rate of 4 000 tpd]. BBE has now been commissioned by SRK Consulting

[Vancouver] to carry out a ventilation/refrigeration study for Sub-Level Retreat [SLR] miningmethod with no backfill to -979 Level for Rory’s Knoll with a production rate of 5 000 tpd. Therevised underground plan is based on truck haulage by ramp with no shaft. BBE was requested toreview the ventilation model and assess the cooling requirements.

The mine plan involves mining down to a limit of -979 Level. Cooled air will be delivered



underground in a dedicated fresh air raise [FAR] from surface. Exhaust from the mine will besimilar to the original feasibility study – two parallel raises with exhaust fans on surface.

The specific scope of work relates to: Review and comment on the ventilation concept and model Determine specifications and costs [CAPEX and OPEX] for the bulk air cooling system Determine specifications and costs [CAPEX and OPEX] for the main surface ventilation fans.

The work related to a feasibility-level of detail and specifically excluded the following: Detailed design drawings for construction, P+IDs, civil rebar details

Detailed engineering specification of equipment For construction design/specifications Health and Safety Codes of Practice EPCM involvement

Sit i it b BBE i


2 DESIGN CRITERIA

The purpose of this section is to outline the key ventilation and refrigeration design parameters tobe used for the feasibility study; the content of this document has been configured in line withGuyana Goldfields Inc. standards as well as generally accepted industry standards. The project isassumed to have the following features and design assumptions:

Total production 5 000 tpd ore600 tpd waste

Maximum depth of mining 979 mbc

2 1 Production Profile and Mine Layout



2.1 Production Profile and Mine Layout

Figure 1 shows the layout of the mine and the planned ramp system. This is the basis of the

VUMA-network model.

Aurora Project Guyana, Rory’s Knoll, Feasibility Ventilation and Refrigeration Study, BBE report 6912 December 2012,

A preliminary production schedule was supplied by the client and is shown in Table 1.

Table 1 Preliminary Production Schedule




2.2 Ore Extraction Method

The SLR mining method is planned to extract rock from the stopes with no backfilling of the voidsallowing the wall rocks to cave in to the extracted stope after all the ore has been removed. Thestope is then sealed to prevent access. There is a 25 m vertical difference between sub-levels and100 m vertical difference between levels. Four sub-levels will be operated on to produce5 000 t/day and one sub-level will be in development.

2.3 Main Airway Veloci ty

Table 2 Main Airway Velocit ies

Return Air Raises [RAR] no personnel 18-22 m/s [1] F h Ai R i [FAR] l 18 22 /



Fresh Air Raises [FAR] no personnel 18-22 m/s [1] Intake airways [personnel] 4-6 m/s

Return airways [personnel and tramming] 6-8 m/sDedicated return airways – no personnel 8-15 m/sDeclines, ramps 6-8 m/sNote1. Pressure drops, economics and practical issues will be taken into account when determining final velocity for

dedicated vent raises and airways.

2.4 Friction Factors

Table 3 Main Airway Friction Factors

Blasted airways [intake/return] [Irregular sides] 0.011 Ns2/m

Galvanised vent ducting / Layflat ducting 0.002 Ns2/m


2.7 Diesel Equipment

Table 6 Final Diesel Fleet

Underground Diesel FleetUnits

RequiredRatedKW

Utilisation kW

Sandvik DD421 Jumbo Drill 2 110 15% 33

Sandvik DS311 Bolter 2 110 15% 33Sandvik DL421-7 Production Drill 2 110 10% 22

Cubex Ares w/ V30 1 130 10% 13

Sandvik LH517 Loader - 17t 5 298 60% 894

Normet Charmec 1610B 2 110 33% 73

CAT IT930 Integrated Tool Carrier 4 116 35% 162



g

CAT AD55B Truck 13 579 60% 4 516

Grader 1 129 60% 77

Boom Truck 1 103 25% 26

Forklift 1 54 25% 14

Toyota LV 20 151 20% 604

Toyota Personnel Carrier 4 151 15% 91

Service Truck 2 110 25% 55

Fuel/Lube Truck 1 103 25% 26

Total 6 639

Ventilation rate for Diesel [0.06 m³/s per kW] 398

Contingency - 15% 60

Rory's Knoll Ventilation Estimate [m³/s] 458


Based on the above relationship reference VRTs are:

79 m Level 24.7 °C154 m Level 25.6 °C229 m Level 26.4 °C304 m Level 27.3 °C379 m Level 28.2 °C

454 m Level 29.1 °C529 m Level 30.0 °C604 m Level 30.8 °C679 m Level 31.7 °C754 m Level 32.6 °C829 m Level 33.5 °C904 L l 34 3 °C



904 m Level 34.3 °C979 m Level 35.2 °C

Table 7 Rock Properties

RockThermal

Diffusivity[m²/s]

Thermalconductivity

[W/m°C]

SpecificHeat

[J/kg °C]

BulkDensity[kg/m3]

Orebody 1.19 x 10-6 3.5 1 041 2 820

Non-Orebody rock hosting mine workings 1.51 x 10-6 4.3 1 027 2 780

Rock thermal properties were measured as part of the previous project. Readings of rock densitiesof 2 530 kg/m³, specific heat of 1 150 J/kg °C and thermal conductivity of 3.4 W/m °C for the orebody were obtained. Subsequently, updated properties have been received, in the above table,

hi h d f thi t d


2.13 Refuge Chambers

Refuge chambers will initially be located at station/landing breakaways, after which they will be re-established on every level to ensure personnel travel no further than 1 000 m from the entrance tothe production section/development heading. This distance is based on a maximum walking timeof 20 minutes, allowing for some horizontal and some up- or down-ramp walking with normal body-worn equipment.

2.14 Second Outlets

In the current mine layout there are no dedicated escape-ways between surface and –979 Level. Itis suggested that if no dedicated escape-ways are planned, then the FARs between surface and -979 Level are equipped with escape ladders to serve the purpose of a secondary egress.Demarcated escape haulages will be maintained between the production zones and the ramp and



p g p pwill be used as a second egress from the mine. In virgin development, where it is impractical tohave these connections, the number of persons working in these areas will be restricted to 50persons and the necessary control measures introduced to minimise the risk.

2.15 Airborne Pollutants [General Air]

All exposures to airborne pollutants are to be within the Occupational Exposure Limits [OEL].

Carbon Monoxide 25 ppm [TWA] Carbon Dioxide 5 000 ppm [TWA]

[NO] 25 ppm [TWA] [NO2] 3 ppm [TWA] Total dust 10 mg/m³ Flammable gas < 1.0%


3 HEAT LOAD AND VENTILATION NETWORK ANALYSIS

A feature of the high-level design process has been a full interactive computer simulation of theheat flow, ventilation and cooling systems to determine air temperatures, flow rates, heat loads andcooling requirements. VUMA3D-network software was used for this purpose. To understand theultimate heat load, simulation models were developed for a number of snap-shots.

3.1 Critical Snap-Shots

A number of critical snap-shots during the build-up of the mine were identified to determine thephase-in of the ventilation / cooling equipment. Maximum depth of mining before refrigeration is required Depth where two refrigeration machines will be required [operating at part load]

Ulti t til ti / li i t



Ultimate ventilation / cooling requirement.

3.2 Heat Load Components

The overall heat load can be broken down into the following main components:

Surrounding rock creates a heat load on the intake system by conduction from the hot virgin rockinterior to the rock surface and into the excavations. This heat component increases with depth asboth virgin rock temperature [VRT] and the extent of exposed rock surface area grows. This heatload also depends on the excavation age [newly mined rock surfaces transmit heat rapidly but this

slows as the rock cools].

Broken rock creates a heat load because it enters the underground environment at a warmtemperature and gets cooled by the air as it leaves the mine. The temperature to which it is cooleddepends on the actual time taken to remove it from the mine but the modelling generally assumes


3.3 Results of VUMA Models

Primary ventilation and cooling requirements indicated by VUMA-network analyses are providedbelow for the critical snap-shots.

3.3.1 Maximum depth of mining before refrigeration is required – Year 3 onwards

Primary Air Requirements

For this scenario it is assumed that uncooled fresh air will be delivered to the production zones.For this phase the rock production is 5 000 tpd and the main diesel fleet consists of eight truckshauling rock to surface, and one LHD per production stope [5-off in total]. Reject air will be drawnup the upcast raise by the main fans located on surface. The minimum air quantity required fordilution of diesel fumes from the diesel equipment is 336 m³/s, as per the diesel fleet requirementsfor Years 3 and 4. However, due to the air requirement in the production zones, the full 460 m³/s is



q prequired. At this flow rate, the design velocities in the ramp and FAR are at the high end of thedesign velocities. A screenshot for this scenario shows the wet-bulb temperature variation on -354Level. In the figure blue indicates cooler zones, while red indicates hot zones.


fed into the ramp at crosscut at -54 Level to provide some cooling in the ramp to keep it below32 °C wb. This will apply from Year 3 of the mining schedule where mining is below -354 Level

3.3.2 Depth where two refrigeration machines will be required – Year 7 onwards

Primary Air RequirementsFresh air will be delivered to the production zones from surface via the FAR with connections from

the FAR to the production and development levels. Reject air will be drawn up the production RARby fans located on surface. The full refrigeration/air cooling system will be installed on surface atthe FAR but will be operating at part load. Minimum air quantity required for dilution of dieselfumes from diesel equipment is 384 m³/s however additional air will be required for the productionzones, so the full 460 m³/s is required.

For this phase the required production is 5 000 tpd from four production sub-levels and the diesel



p q p p pfleet is as shown in Table 6.

The maximum face temperature is predicted to be 28.3 °C wb which occurs on -604 Level.


3.3.3 Ultimate ventilation / coo ling requirement

Primary Air RequirementsFresh air will be delivered to the production zones from surface via the FAR with connections fromthe FAR to the production levels from -879 Level to the maximum depth of mine. Reject air will bedrawn up the production RAR by fans located on surface. The full air cooling system of 11.0 MWRwill be implemented on surface at the FAR. Minimum air quantity required for dilution of diesel

fumes from diesel equipment is 460 m³/s.

For this phase the required production is 5 000 tpd from five production levels and the diesel fleetis as shown in Table 6.

The maximum production face temperature is predicted to be 28.4 °C wb which occurs on -954 Level. The ramp temperature is predicted to be 31.9 °C wb.




4 SUMMARY STATEMENT OF ULTIMATE PRIMARY VENTILATION AND AIR COOLERREQUIREMENTS

For reference purposes, and as an overall summary statement, the ultimate needs are as follows[phase-in considerations are presented in Section 9]:

Total ventilation Total air flow specification [D/C and U/C flow] 512 kg/s

Downcast ventilation Intake ramp 212 kg/s Fresh Air Raise 300 kg/s



g

Upcast ventilation 4.0 m diameter upcast ventilation raise 256 kg/s 4.0 m diameter upcast ventilation raise 256 kg/s

Ai r coo lers Main downcast raise surface air cooler duty 11.0 MWR

It should be noted that the ventilation and cooling refrigeration system is sized for miningto -979 Level in this study. This includes the sizing of the infrastructure on surface, refrigerationmachines and cooling towers. The previous study examined the cooling requirements downto -1 230 Level and required more cooling to achieve acceptable conditions.


5 PRIMARY VENTILATION

5.1 Intake infrastructure

Fresh air will enter the mine via the FAR [4.0 m diameter] and the ramp [5.5 m x 6.0 m]. A bulk aircooler will be installed on surface at the FAR. Uncooled air will flow in the ramp, which will providefor diesel dilution in the ramp. Some cool air will enter the ramp on -54 Level via a crosscutconnecting FAR to the ramp. Cross-connections between the FAR and the production zones occuron every level from -79 Level.

Air velocities in the ramp are at the high end of the design criteria at 8.0 m/s maximum whichoccurs directly after cooled air from the FAR is supplemented into the ramp for additional cooling.The velocities in the FAR are 19.9 m/s at maximum and decrease to 13.2 m/s after air is

l t d t th f th FAR



supplemented to the ramp from the FAR.

5.2 Return Infrastructure

There are two primary return air raises [RAR], one to return air from the production sub-levels andone to return air from the decline and ramp. Both RARs are 4.0 m diameter raise bore holes [RBH].The production RAR is raise bored in stages from the bottom of the mine [-979Level] to -79 Level and then to surface. The decline RAR is raise bored in stages from -879 Levelto -79 Level and then to surface. The main fans will be installed on surface. Each productionsub-level will breakthrough directly into the production RAR. The decline RAR will return supply air

from the ramp to surface. Breakthroughs into the RARs will be every 25 vertical meters for theproduction RAR and every 100 vertical meters for the decline RAR.

It is recommended that the two RARs are connected to each other to balance the return air flow atthe top of the RARs and as lo as possible in the mine [at the bottom of the decline RAR at 879


6 SECONDARY VENTILATION AND VENTILATION CONTROL ISSUES

6.1 Ramp Development

The ramp development will be ventilated by a forced fresh air supply ducted in 1 000 m segments.The exhaust air will free flow back to the ramp until the first breakthrough to the ramp RAR isreached. From there the air will be exhausted via the RAR. The duct will then move to develop thesecond segment of the decline. The whole process will be repeated until the entire ramp is fullydeveloped. For the ramp development when temperatures will exceed the set limitations some ofthe cool air will be ducted to the face from the sublevel ventilation circuit FAR. Sufficient air will beprovided to the face zone for one LHD and one truck.

The sublevel drives development will be ventilated either as the ramp extension or a branch off,til th t f til ti i i l t d ith th bl l b kth h



until the system of ventilation raises is completed with the sublevels breakthroughs.

6.2 Service Ventilation

Pump stations, sub-stations and other service ventilation centres will be ventilated in series withthe fresh air delivered to the production zones. Maintenance workshops will be ventilated inparallel to the fresh air and returned directly to the RAR.

6.3 Product ion Zone Ventilation

In determining the minimum ventilation requirements for stoping zones, it has been assumed thattrucks will only operate in the ramp and on sub-levels between the ramp and re-muck bays locatedclose to the ramp. The total ventilation flow in the FAR will be sufficient to meet the dilution criteriain the production zones and thus no provision has been made to supply additional air to a sub-levelfor dilution of diesel pollution from trucks


Auxiliary force fan 90 kW

6.4 Regulators to Upcast Ventilation Raises

Ventilation regulators will be provided at the return air raise connection on each of the 37 sub-levels. Typically these would be roll-up garage-type doors, adjustable louvers or guillotine-typeregulators operated by manual chain block, see Figure 5. Alternatively it is possible to use a

ventilation duct in a permanent seal for adjusting the flow of air. The flow range over which flow isto be regulated is from 0 m³/s to about 40 m³/s.

The regulator will be closed on non-producing levels but would be opened to varying degrees onproducing levels. As mining advances deeper, then more air will be required on a level for heatdilution and the regulators will be used to ensure that too much air is not allocated to upper levelsto the detriment of the lower hotter levels A ventilation technician will monitor underground



to the detriment of the lower hotter levels. A ventilation technician will monitor undergroundventilation [at least three times per week] and flow sensors and air quality sensors will report to

output on surface. The ventilation technician will operate the regulators, and would leave writteninstructions in the shift boss daily log book. Shift bosses will have the authority to adjust ventregulators. The central control centre will monitor actual air flow and diesel equipment reporting toa level.


Combined CO/NO2/Smoke Returns from all working crosscuts Main RAW on -79 Level Main fans

Ai r Veloc ity and Temperature Monitors Main intake levels Main RAW on -79 Level Main fans

Telemetry System A telemetry system coupled to the early warning devices is required. Measurements must betaken in ‘Real time’ and contin o sl rela ed to a central control room Appropriate alarm le els



taken in ‘Real time’ and continuously relayed to a central control room. Appropriate alarm levelsare to be set that will allow control room personnel to promptly warn employees of potential life

threatening occurrences [such as irrespirable atmosphere]. A stench gas will be released into themine air to warn personnel to leave the work areas. The early warning devices will also provide anaudible alarm should an ‘Alarm’ level be reached.

6.6 Refuge Bays

Refuge bays will initially be located every 1 000 m along the main ramp and levels to ensurepersonnel travel no further than 1 000 m from the furthest face/development heading. Where

practical refuge bays will utilise redundant excavations, where these are unavailable, suitableexcavations will be blasted or portable refuge bays set-up. General layout is shown in Figure 6.This equipment will be part of the cost estimate as reflected in Section 10.


Refuge Chambers

570mm Ventilationpipe with stretcher and

blankets

FluorescentLights

Ai r and Watersupply

Identification numberpainted inside and outside

Drinkingwater

Two Compressedair sil encers

100mm Compressed

air pipe

Toiletwith Lid



blankets

Drain

ConspicuousLight

Notice Board

Mandooropen inside

Note!1) Notice boards will display treatment procedures forelectric shock, heatstroke, gassing and drowning.2) Refuge bay Operating procedure.3) Ventilation Department will stipu late size, positionand layout of bay.4) No working valves may be installed between maintake-off and refuge bay.5) Refuge bay will be positioned a maximum of1000m from the workplace.

Telephone


7 FAN STATIONS

7.1 Main Fan Station Specifications

7.1.1 Aerodynamic performance c ri ter iaThe peak required aerodynamic performance criteria for the main fan stations for Rory’s Knoll is asfollows:

Number of fans installed 2 off Number of fans operating 2 off Total maximum required volumetric flow rate per fan 230 m3/s [487 000 cfm] Maximum static pressure differential 3 400 Pa [13.7” wg] Maximum required mass flow rate per fan 256 kg/s Rated motor power 1 000 kW [1 350 hp] Inlet density to fan 1 10 kg/m3



Inlet density to fan 1.10 kg/m Temperature 26 °C saturated

Note: Above pressure includes system pressure and fan losses.

At the early stages in the mine life when the pressure drop through the mine is less than theultimate condition, the main fans blade angles can be adjusted to accommodate the lower pressureand consume less energy.

A typical fan curve is shown in Figure 7.





The main fan station at Rory’s Knoll mine will consist of an independent axial flow fan station ateach main upcast raise borehole. Refer to drawing BBE-12088-04-001-01 for a general layout ofthe fan station.

The specifications for the main fan stations are as follows: Aerofoil impeller diameter 2 800 mm Fan speed 890 rpm Motor rated power 1 000 kW

Each axial fan station will include the following: Inlet shaft top bend 4 m dia. Shaft safety-screen 4 m dia. Access hatches



Access hatches Flexible connectors Self-closing doors Transition ducting Maintenance platform Single inlet ducting 2 800 mm diameter axial fan c/w steel impeller hub, aluminium blades, inlet cone, dome,

internal fairing, fan casing, guide vane casing, external motor base frame, turbo flex couplingand bearings

Floating Carden shaft Holding brake Set of guards Fan diffuser Discharge bend c/w turning vanes


7.1.6 Fan noiseTo reduce fan noise provision is made for outlet silencers, the fan station will have a verticaldischarge to direct fan noise upwards. The initial indication is that an inlet silencer will not berequired.

7.1.7 Construction and labour requirementsThe predicted construction period for the main ventilation fan stations is as follows:

Civil works 3 MonthsMechanical and Electrical Works 2 MonthsCommissioning 1 Month

These time frames are assuming that all the equipment has been delivered and is on site ready forinstallation.



The personnel requirements are as follows:

Civil works 20 PeopleMechanical Works 10 PeopleElectrical Works 5 PeopleCommissioning 5 People

7.1.8 GeneralProvision must be made for electrical power supply to the sites, monitoring systems to mine controlcentre, roads, security fencing, etc.

Although standby fan-motor sets will not be required provision should be made for strategic sparesas given below:


Each fan will include the following main components:

Fan-unit IP55 enclosure Inlet and discharge axial silencers Instrumentation and local hard wiring and switches for temperatures and vibration One complete spare fan-motor set, commissioning and operational spares for two years.





7.3 Product ion Zone Fan Specifications

For each production crosscut, the following is required: Number of fans installed 1 off Number of fans operating 1 off Total maximum required volumetric flow rate per fan 35 m3/s [74 100 cfm] Maximum static pressure differential 1 500 Pa [6.8” wg]

Maximum required mass flow rate per fan 40 kg/s Rated Motor Power 90 kW [120 hp] Inlet density to fan 1.24 kg/m3 Temperature 27.0/30.0 °C wb/db Duct diameter 1 300 mm minimum

Each fan will include the following main components:



Fan-unit. VFD drive units [IP55 enclosure or better] with variable frequency control. Inlet and discharge axial silencers. Instrumentation and local hard wiring and switches for temperatures and vibration.








The BAC will have the following process specifications [summer day 14h00]:

Air cooler air duty 11 MWR Air cooler air flow 300 kg/s Air temperature onto air cooler [ambient air] 26.0/31.4 °C, 101 kPa Air temperature off air cooler 15.7 °C wb

8.3 Bulk air cooler Fans

Inlet air cooler fans [2 off] will force air through the air cooler and down the FAR. The air pressuredrop across the actual air cooler will be modest but the fans will be required to force air flowthrough the cooler and into the FAR and must provide for the pressure drop requirements of the aircooler and bend/entry into the FAR.

Typically the fans will be 2 4 m diameter directly driven axial flow units installed with inlet cone and



Typically the fans will be 2.4 m diameter, directly-driven axial flow units installed with inlet cone andsafety screen, silencers, self-closing doors, fan-motor sets, flexible duct couplings, outlet diffuserducting and support structures. Fan blade angles will be manually adjustable and the fans willhave variable speed drives. Noise level will not exceed 81 dBa at 6 m from the fans.

The fans will be relatively low pressure units driven by direct-coupled in-line 160 kW, 6 pole motorswith the following general specification:

Number of fan units 2 off Air mass flow per fan unit 150 kg/s

Air volume flow per fan unit 133 m³/s Suction density 1.14 kg/m3 System total pressure nominal 0.5 kPa [at outlet from diffuser] Fan diameter 2 400 mm


Scope of supply will include initial charges of refrigerant [commissioning requirements to be listedand costs indicated separately].

Lubrication oil will be force-fed to all compressor bearings and rotating surfaces from an externaloil pump and reservoir. Dual oil filters of replaceable cartridge type equipped with service valveswill be provided.

Oil cooling will be achieved by a refrigerant-oil-cooler. Automatic oil return/recovery system fromevaporator is included. Complete lubrication system [piping and components] will be factoryassembled and tested at the manufacturer’s works.

The condensers and evaporators will be shell-and-tube type with water passing through the tubesand refrigerant on the shell side. The shells will be manufactured from carbon steel and painted



with high specification epoxy coating. The tubes will be 90/10 CuNi [with clad tube sheets in same

material] with high efficiency enhanced internal and external surfaces. Chilled water within therefrigeration system will remain separate and independent of the mine service water system.Thermal insulation will be applied to the surface of all vessels, piping, flanges, valves and fittingsthat have potential for condensation. The insulation will be covered by galvanised steel sheetingfor mechanical protection.

The system will also include a refrigerant leak detection system and a refrigerant pump-downfacility that is permanently hard-piped to the chillers. Suitable access platforms for the refrigerationmachine and other equipment will be an integral part of the system. The final refrigeration systemdesign parameters are as follows:

Total system cooling effect 11.0 MWR Process duty of refrigeration machines 11.6 MWR


8.7 Water and Pump Systems

The condenser water system will simply circulate water from the CCT sumps to the plants andback. Provision will be made in the costing for 2 off [plus 1 standby] pump-motor sets of 110 kWrating for this purpose.

The evaporator water system will simply circulate water from the BAC sump to the plants.Provision will be made in the costing for 1 off [plus1 standby] pump-motor set of 160 kW rating forthis purpose.

The re-spray pumps will spray the water in the spray chamber. Provision will be made in thecosting for 1 off pump-motor set of 110 kW rating for this purpose.

These pumps will all be horizontal split casing, single-stage, centrifugal units directly coupled tofour pole motors.



four pole motors.

Each pump system will be protected by strainer screens installed in sumps. There will beautomatic water strainers installed on both the condenser and evaporator circuits.

8.8 Refrigeration System Preliminary Motor List

The main motors [nameplate rating] in the refrigeration system will be as follows:

Table 8 Main Motor List for Surface Refrigeration System

Main Motor Summary

No.Installed

Voltage[3 phase]

Ratedeach

Ratedtotal

Typicalabsorbed,

summers day


8.9 Construction and Labour Requirements

The predicted construction period for the BAC, refrigeration plant rooms and cooling tower is asfollows:

Civil works 8 MonthsMechanical and Electrical Works 4 MonthsCommissioning 1 Month

These time frames are assuming that all the equipment has been delivered and is on site ready forinstallation.

The personnel requirements are as follows:

Civil works 100 People



Civil works 100 PeopleMechanical Works 25 People

Electrical Works 25 People

8.10 Absorption Chillers move to end

The use of absorption chillers was investigated as a possible alternative to conventional chillers.Since all the plant electrical power is generated from diesel generators, the exhaust gas could beutilised as a free heat source for the powering of absorption chillers.

Absorption chillers operate at much lower efficiencies than conventional refrigeration machines.This results in a much higher capital cost for an absorption chiller installation to achieve the samerefrigeration. The use of absorption chillers will therefore be justified if the saving in energy costswould outweigh the difference in capital cost.


8.10.1 Absorption Chi ller Select ion

Various arrangements and options for absorption chillers were investigated and the mosteconomical and efficient solution was found to be the Broad absorption chillers. This chiller isconnected directly to the generator exhaust and does not need an additional exhaust to hot waterheat exchanger as required for the bulk of absorption chillers found in the industry.

The model suitable for this application is the Broad BDE100 absorption chiller. This model chillerhas a refrigeration capacity of 1.16MW from an exhaust gas flow of 28450kg/hr @ 300°C. This isthe exhaust gas production of the 16V32 Wartsila generator at an electrical load of approximately4.2MW.

It will be possible to install more than one Broad absorption chiller depending on the generator loadand the area available.



8.10.2 Infrastructure, Area Allocation and Utilities

The Broad chiller is supplied as a complete package and can be ordered with a metal enclosure,which make it possible to install the chiller outside, thus no building would be required. Each chilleralso comes with its own cooling tower and exhaust gas bypass valve.

This chiller needs to be installed as close as possible to the generator exhaust. The cooling watercan be pumped through insulated pipes to the refrigeration plant building and bulk air cooler. Theopen area required for each chiller package is 6m x 11.5m. In addition, an area of 6m x 4.5m isrequired for each chiller’s cooling tower and an area of 1.6m x 1.6m is required for locating anexhaust bypass 3-way valve.


Table 9 Absorption Chiller Capital Costs

Guyana Gold Aurora Project

Rory's Knoll SLR Mining Method Base date: 12-Dec-12

ABSORPTION CHILLER PACKAGE FOR 1.16MW COOLING

FEASIBILTY COST ESTIMATE Rev. date: 12-Dec-12

Summary Cost

1 BDE100 chiller 199 215$

2 Condenser and evaporator pumps 79 800$

3 Metal Enclosure 14 440$

4 Cooling tower 24 220$



The above costing excludes the electrical supply feeders and utility connections The piping

g

5 Exhaust valve (DN900) 48 390$

6 Commissioning + Expenses 50 000$7 Local transport 30 000$

8 Piping infrastructure, installation & electrical. 300 000$

Subtotal directs 746 065$

EPCM, project services and indirects 25% of directs 186 516$

Contingency 10% of directs 74 607$

Total 1 007 188$




Table 11 Capital Cost Payback – No Escalation

Scenario: Capital cost vs accumulated energy costs calculated with no escalation

Year

Capital cost

[US$]

Absorbed

power saving

[kW]

Energy costs per

kWh [US$]

Energy Saving

per hour

[US$]

Energy saving

per year (US$)

Accumulative energy

cost saving [US$]

1 0.00 0.00 0.265 0.00 0.00 0.002 0.00 0.00 0.265 0.00 0.00 0.00

3 0.00 0.00 0.265 0.00 0.00 0.00

4 0.00 0.00 0.265 0.00 0.00 0.00

5 0.00 0.00 0.265 0.00 0.00 0.00

6 1 007 187.75 108.44 0.265 0.00 0.00 0.00

7 1 007 187.75 108.44 0.265 0.00 0.00 0.00

8 1 00 18 108 44 0 26 28 4 201 394 91 201 394 91



8 1 007 187.75 108.44 0.265 28.74 201 394.91 201 394.91

9 1 007

187.75 108.44 0.265 28.74 201

394.91 402

789.82

10 1 007 187.75 108.44 0.265 28.74 201 394.91 604 184.73

11 1 007 187.75 108.44 0.265 28.74 201 394.91 805 579.64

12 1 007 187.75 108.44 0.265 28.74 201 394.91 1 006 974.55

13 1 007 187.75 108.44 0.265 28.74 201 394.91 1 208 369.46

14 1 007 187.75 108.44 0.265 28.74 201 394.91 1 409 764.37

15 1 007 187.75 108.44 0.265 28.74 201 394.91 1 611 159.28


fibreglass condenser cooling tower, installed in close proximity to the generators. This modeland arrangement was found to be the most economical.

No deductions in capital costs were made should it be decided that the conventionalrefrigeration plant and equipment be reduced.

The 5% escalation used in scenario 1 was based on assumptions and was done to show theeffect of such an escalation. Other scenarios and escalation values can be investigated ifrequired.

8.10.5 Absorption Chi ller Investigat ion: Conclus ion

It would be possible to achieve a net saving with the installation of absorption chillers over the lifeof mine. Payback should be achieved within five years from time of installation. The analysisassumed that the originally specified standard refrigeration chillers would be purchased but run at



g y p g ppart load while the absorption chiller is operating. Further studies could examine the possibility of

additional savings by reducing the standard chiller size.

Should the mine wish to pursue this option further, details of the generator installation andavailable area will have to be investigated and a more detailed costing should be done.


9 PHASE-IN OF CAPEX FOR MAIN COMPONENTS AND LIFE-OF-MINE POWERREQUIREMENTS

Phase-in of the main components for CAPEX estimates are discussed below. No rampdevelopment schedule was available but the preliminary production build-up in Table 1 indicatesthat full production is achieved by Year 2 after start of production.

9.1 Main Fans

As soon as the first of the upcast raises is completed, the first of the upcast fan stations must beavailable. The first fan will ventilate the ramp-up of production as well as the part of the declinedevelopment. The second fan station will be commissioned on completion of the second of theupcast raises.



9.2 Refrigeration and Cooling

Refrigeration / air cooling will be required as soon as mining extends below -354 Level. This is thecase whether the development of the raise is delayed or the raises developed concurrently with theramp development.

9.3 Power Profiles

Estimated primary absorbed fan power and refrigeration motor power for the critical years areprovided in Table 12. Secondary absorbed fan power will be approximately 0.45 MW during

development.

Figure 13 shows the predicted utilisation of the ventilation fans and the refrigeration system as themine ramps up to full production.


Using this profile the predicted power usage and operating costs are detailed in Table 12. Anelectricity cost of $0.265/kWhr was used in the operating cost calculations.

Table 12 Power Requirements of Main Components and Operating Costs

Components Year -2 Year 3 Year 7 Year 11 Year 15

VentilationPrimary ventilation 357 kW 1 242 kW 1 581 kW 1 770 kW 1 844 kW

Development fans 180 kW 180 kW 180 kW 180 kW -

Production fans 110 kW 810 kW 810 kW 810 kW 810 kW

Ventilation Total 647 kW 2 232 kW 2 571 kW 2 760 kW 2 654 kW

R f i ti S t 1 005 kW 1 916 kW 2 692 kW 2 692 kW



Refrigeration System - 1 005 kW 1 916 kW 2 692 kW 2 692 kW

Total 647 kW 3 237 kW 4 487 kW 5 452 kW 5 346 kW

Annual Operating Cost* $1.5M $7.5M $9.2M $12.0M $11.8M

Annual Maintenance Cost $1.1M $0.9M $0.4M $0.4M $0.2M

Total OPEX Cost* $2.6M $8.4M $10.5M $12.4M $12.0M

Note

*Operating cost assumes continuous operation for 365 days per year This will be the maximum potential


10 CAPITAL ESTIMATE OF MAIN COMPONENTS

10.1 Surface Main Fan Station[s]

The estimate is based on extrapolation of costs from BBE archives for similar fan stations and

budget costs from Actom. Capital cost breakdown and vendor quotes are given in Appendix 1

COST ESTIMATE SHEETS. CAPEX provision for two fan stations will be $6.8M.

Includes : Main fan/motor sets, shaft bends, transition pieces, self-closing doors, duct work, civilwork, electrical work, strategic spares, etc.

10.2 Secondary Ventilation Equipment

The estimate for the secondary ventilation equipment is based on specific quotes for individual



The estimate for the secondary ventilation equipment is based on specific quotes for individual

items [Nov 2011] and information from BBE archives. Capital cost sheet for secondary ventilationequipment is given in Appendix 1

COST ESTIMATE SHEETS, value is estimated at $8.5M.

Includes : Auxiliary ventilation requirements [ducts, fans, air locks], refuge bays, ventilationstoppings and personal self-contained self-rescuers.

10.3 Surface Cooling

The estimate is based on extrapolation of costs from BBE archives and cost estimates provided

from Johnson Control International - York. The capital cost breakdown of the refrigeration project


Table 13 Ventilation Officials’ Instruments

Instrument Number offCost each

$Total

$

1 Davis anemometer in carry case 3 1 142 3 426

2 Kestrel 4000 anemometer 3 575 1 725

3 Laser distance meter 3 485 1 455

4 ALNOR micromamometer [measure pressure to 5 kPa] 3 1 184 3 552

5 Greisinger GPB3300 Barometer 3 262 7866 Haden double limb whirling hygrometer & case 6 45 270

7 Hygrometer thermometers -5 to 50 deg [0.5 grad] 8 14 112

8 2kg aluminium dust 2 122 244

9 Puffer bottles 10 1 10

10 Anemometer extension rods [4 off 0.5 m lengths] 12 18 216

11 Wet strength books 10 16 160

12 Smoke tube kit 3 122 366



12 Smoke tube kit 3 122 366

13 Spare Kestrel vanes 3 35 105

14 Stop watch 3 66 198

15 Instrument carry bag 3 68 204

17 Sidepak Personal Aerosol monitor AM510 2 7 020 14 040

16 CEL-320S Cassella Type 2 Impulse integrating sound level meter 1 1 717 1 717

17 CEL-110 Type 2 Acoustic Calibrator 1 739 739

29 325

10.5 Early Warning System

Early warning system includes monitoring instruments, alarms and communication system,itemised in Table 14. Estimate based on BBE archives. This cost is included in SecondaryVentilation costs [Section 10.3].

I l d d CO bi d CO/NO / k St h G t t t d i l it


10.6 Estimate of Freight Requirements

10.6.1 Refrigeration plant and BAC

Table 15 below details the civil works materials transport requirements and costs to site.

Table 15 Civils

ComponentMass

[tonnes]No. of

containersFreight to

Georgetown

Road Transportto Aurora

@ $0.012/lb

Cement [powder] 465

Steel [reinforcing] 110

Formwork 45



Scaffolding 60

Total 680 35 $157 500 $17 990

Table 16 below details the transport requirements and costs to site of the refrigeration machinesand pumps.

Table 16 Refrigeration Machines

Component Mass[tonnes] No. ofcontainers Freight toGeorgetown

Road Transport

to Aurora@ $0.012/lb

Refrigeration machines [2 off] 546 off, 40 ft flatrack [6 off 9


10.6.2 Main fans and secondary ventilation

Table 18 below details the transport requirements and costs to site of the main and secondary

fans.

Table 18 Main Fans and Secondary Ventilation

ComponentMass

[tonnes]No. of

containersFreight to

Georgetown

Road Transportto Aurora

@ $0.012/lbFan Station [2 off] Ductsteelwork, electricals

2 x 24.164 2

Cement for civils420 x 50 kg

= 211



21

Main fans [2 off] 2 x 13.75 1Ventilation ducts 22.43 1

Main ramp development fan,Strategic spares for main fans

25 1

Ventilation instruments, earlywarning system

22 1

Total 146 7 $31 500 $3 860


11 POTENTIAL RISKS

During the course of the project the following major risks were identified. Hot ground water entering mine workings High temperatures in lower ramp Vehicle fire

Additional equipment High air velocities in lower FAR for secondary escape means.

11.1 Hot Ground Water entering Mine Workings

The assumption is that the ground water filters through the rock from surface and arrives in theworkings at a temperature lower than the local VRT. The ground water quantity used in this studyis significantly less than that used in the previous studies It must be noted that should significant



is significantly less than that used in the previous studies. It must be noted that should significant

quantities of ground water be experienced at temperatures close to the VRT, additionalrefrigeration may be necessary to provide acceptable temperature conditions underground.

The quantity and temperature of ground water entering the mine must be monitored from the startof development. If inflows at rates different to the current assumptions are encountered, thepotential effect on workplace temperatures must be simulated and mitigating steps taken. Groundwater should not be allowed to flow freely on the footwall. It should rather be piped back to themain pumps in [plastic] piping to avoid contact with the air.

11.2 High Temperatures in Lower Ramp

In the lower levels of the mine, a wet-bulb temperature of 32 °C wb is predicted in the ramp. Thetrucks in the ramp are to be provided with enclosed air-conditioned cabins This is a risk since


Diesel has recently been classified as a carcinogen by the International Agency for Research onCancer [IARC]. There is potential with this classification as a carcinogen that air quantitiesrequired for diesel dilution could be increased in future.

11.5 High Air Velocities in Lower FAR for Secondary Escape Means

At the highest point in the FAR before air is delivered to the production zones, there are high

velocities in the FAR of 12.5 m/s. This could be a problem in the event that the FAR is required tobe used as a secondary means of escape in the event of an evacuation. This can be addressedby installing the escape ladders in an independent duct in the FAR or by lowering the velocity inthe FAR by reducing the fan speed in the event of an evacuation.




12 OPPORTUNITIES

12.1 Optimum Airway Sizing

There is an opportunity to further reduce capital and operational costs by optimizing the size of theFARs and RARs. A previous study [Aurora Project Life-of-Mine Feasibility Study Optimization, BBEreport 6012, September 2012] was conducted on the optimization of the FAR and RAR sizes. It

was concluded that larger raise boreholes of 5.0 m in diameter are more economical than thecurrent sizes of 4.0 m diameter. A copy of this study is included in Appendix 2.

12.2 Additional Absorption Chillers

Providing there is sufficient area close to the generators on site and sufficient waste heat available,there is an opportunity to use multiple absorption chillers to reduce the requirement for



pp y p p qconventional chillers.

13 ACKNOWLEDGEMENT

This work has been carried out in close collaboration with personnel from SRK in Vancouver.

R HoodBBE Consulting


APPENDIX 1

COST ESTIMATE SHEETS

Refrigeration Plant and Air Cooler




3 Air cooler mechanicals and internals

3.1 Spray header and piping 152 000$

3.2 Mist eliminator 97 000$

3.3 Air cooler fans, non-return doors and silencers 467 500$

3.5 Strategic spares provision 48 000$

3.6 Installation and commissioning [including P+Gs] 229 000$

3.7 Items not measured 10% 99 000$

Subtotal 1 092 500$

Excludes air cooler civils, connection to shaft, system instrumentation and electricals.

4 Pumps, valves and piping

Includes 2-off axial fans with motors, silencers, diffusers and non-return doors, internal air cooler piping, nozzles,

mist eliminators and mist eliminator support frame work.



4.1 Evaporator, condenser and re-spray pumps 196 000$

4.2 External air cooler, condenser and service piping and valves 413 500$4.3 Insulation 87 000$

4.4 Strategic spares provision 70 000$

4.5 Valve, Pipe and Pump installation contractor 279 000$


Subtotal 1 150 500$

Includes local piping, valves, pumps, motors and insulation of chilled water pipes.

Excludes civils, pipe supports, electricals and water supply to site.

5 Electricals

5.1 Switchboard, transformer, MCC, and auxiliary items 609 500$


Main Fan Station [2 of f]




Main Fan Station Vendor Quote

AXIAL FLOW FAN COSTING

BUDGET PRICING ‐ 230m^3/sec @ 3440Pa

Item Description Qty Price Each Price Total

1 Inl et shaft

top

bend

dia.

4m 1 132

933.33$

132

933.33$

2 Shaft Safety screen 4m dia. 1 2 222.27$ 2 222.27$

3 Access hatches 1 8 282.80$ 8 282.80$

4 Flexible connectors 2 3 444.27$ 6 888.53$

5 Fan inlet safety screen 1 1 666.53$ 1 666.53$

6 Self closing doors 1 14 177.73$ 14 177.73$

7 Transition ducting 1 9 611.07$ 9 611.07$

8 Maintenance platform 1 9 611 07$ 9 611 07$



8 Maintenance platform 1 9 611.07$ 9 611.07$

9 Straight inlet

ducting 1 49

722.27$

49

722.27$

10

Fan Axial Type 2800 mm diameter c/w steel impeller hub, aluminium

blades,inlet cone, dome, internal fairing, fan casing, guide vane

casing,external motor base frame, turbo flex coupling and bearings

1

317 523.60$ 317 523.60$

11 Carden shaft, pre‐assembly, bolts and nuts 1 33 535.87$ 33 535.87$

12 Holding Brake 1 9 999.87$ 9 999.87$

13 Motor 1000kW 6P 4.1kV ‐ 60Hz 1 190 000.00$ 190 000.00$

14 Set of guards 1 2 222.27$ 2 222.27$

15 Fan diffuser 1 25

544.27$

25

544.27$

16 Discharge bend c/w turning vanes 1 49 777.60$ 49 777.60$

17 Discharge evase after bend 1 24 917.33$ 24 917.33$

18 Outlet screen 1 1 677 60$ 1 677 60$


Secondary Ventilation Equipment




3 Refuge chambers Unit cost No off

3.1 Supply materials and install 20 000$ 5 100 000$


Subtotal 110 000$

Excludes Excavation costs, power supply to site

4 Ventilation seals and regulators Unit cost No off

4.1 Wall seals 2 000$ 37 74 000$

4.2 Flow regulator 17 000$ 37 629 000$

4.3 Two-door air locks 25 100$ 37 928 700$


Subtotal 1 794 700$

I l d i t ll ti i i

Includes Door, Pipes(air, water), First aid bag/stretcher, Telephone, Valve/Silencer(air), Air whistle, Flash Light,

Construct ion materials



Includes installation provision

Excludes

5 Ventilation instrumentation Unit cost No off

5.1 Gas detection 12 000$ 37 444 000$

5.2 Self contained self rescuers and rack 300$ 200 60 000$

5.3 Ventilation officials instruments 29 000$ 1 29 000$


Subtotal 586 000$

Excludes power supply to site, lamp and control room infrastructures

Included Dual gas meters, functional test points, chager-auto-calibrator






APPENDIX 2

EXAMINATION OF THE SIZES OF THE UPCAST/DOWNCAST VENTILATION AIRWAYS

[Aurora Project Life-of-Mine Feasibility Study Optimization, BBE report 6012, September 2012]

Based on the following parameters, a comparison between different upcast/downcast ventilationshaft diameter sizes is detailed:

Fresh air delivery via 4.0 m diameter Fresh Air Raise [FAR] including refrigeration plant and

bulk air cooler located on surface.

Fresh air delivery via main ramp [uncooled].

Parallel 4.0 m diameter Return Air Raises [RAR] from bottom of mine, main fans located on



Parallel 4.0 m diameter Return Air Raises [RAR] from bottom of mine, main fans located on

surface Small quantity of fresh air delivered in 2.4 m diameter second escape raise.

Deepest level -720m Level.

Truck haulage to surface via ramp.

Nominal diesel fleet 10 082 HP, minimum air for diesel fume dilution 905 000 cfm [427 m³/s],

assuming 90% average utilisation and 100 cfm per HP rated [0.06 m3/s per kW] at point-of-use.

Including contingency, flow estimation by SRK 1 000 000 cfm [472 m³/s]. The total flow could

be more than the minimum required for diesel fume dilution when taking into account heat,leakage and secondary ventilation provision.

Electricity cost $0.30 / kWhr.

A i l i f 3 ³/ [40 k / ]


6 000

8 000

10 000

12 000

f c o s t s [ k $ ]

FAR - Economic airway size



0

2 000

4 000

6 000

7.5 10.0 12.5 15.0 17.5 20.0 22.5 25.0

P r e s e n t v a l u e

o f

Velocity [m/s]


The table below shows the difference in present value costs for two different size shafts for the

FAR [surface to 360 Level] and RAR [270 Level to surface]:

Total PV cost [M$]

Diameter

[m]

UC Shaft

[2-off]DC Shaft

4.0 10.95 9.23

5.0 6.55 5.00

Saving: 4.40 4.23

If the 5.0m diameter shafts are selected, the savings are in the order of $9M when compared to the

4.0m diameter shafts.



Recommendation

It is recommended that the FAR and main RARs are sized at 5.0 m diameter rather than 4.0m

diameter to minimise the velocities in the shafts and hence the total costs of the FAR and RAR.

However, there may be other considerations which mean that the team selects the 4.0m diameter

raise boreholes, for example, additional costs of the 5m diameter raise borehole cutter head.

The current size of 2.4m for the escape way and the 5.5m x 6.0m ramp are acceptable as thevelocity in these are below the maximum velocity of 6.0 m/s.























CONCEPT STUDY OF

THE GUYANA GOLDFIELDS

AURORA PROJECT UNDERGROUND DEWATERING PUMPING SYSTEM

Prepared for:Guyana Goldfields Inc.

TABLE OF CONTENTS

1 INTRODUCTION .. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1

2 SCOPE OF WORK .. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1

3 MINE DEWATERING... . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1

3.1 Water Inflow ........................................................................................ 1

3.2 Pumping System ................................................................................. 2



p g y

3.2.1 Pumping Considerations .......................................................... 2

3.2.2 Pumping Volumes .................................................................... 4

3.2.3 Pump Selection ........................................................................ 4

3.3 Flow Analysis ...................................................................................... 5

3.3.1 Surge Pressure Analysis .......................................................... 5

3.4 Pumping Normal Operations ............................................................... 7

3.5 Flood Pumping Operations .................................................................. 7

3.6 Main Pump Station Pipeline ................................................................ 8

3.7 Pump Stations ..................................................................................... 8

LIST OF APPENDICES

APPENDIX 1 - DRAWINGS

APPENDIX 2 - CAPITAL AND OPERATING COST ESTIMATES



1 INTRODUCTION

Goldfields Inc. is currently undertaking a revision of the Bankable Feasibility Study on their Aurora

Project in Guyana. It is understood that SRK Consulting (Canada) are undertaking portions of this

work. Royal HaskoningDHV (RHDHV) was approached by SRK Consulting to assist with the

underground dewatering study work.



The Aurora Gold Project is situated in the forested region of Guyana, South America,approximately 170 km west of the capital Georgetown.

2 SCOPE OF WORK

The scope of work for the concept level dewatering study includes:

Review of Input data which includes the following:

Mi D i d S h d l



order to cope with these variations the mine dewatering system must be

capable of functioning effectively over a wide range of operating conditions.

The pumping infrastructure must be capable of growing as the mine

develops and the depth increases over time. In the final phase, defined in

this project, the bottom of the mine will be 1,000 m below the surface portal.



For the mining layout considered here the pumping infrastructure required fornormal operation the pumping system will have the following pumping

sections:

Main pumping system capable of pumping the required volume of water

from the mine.

Development pumping system: to collect water from the decline

development and pump to the main pumping system.

Local pumping to collect water from the development face to pump to the

i i t

water be of high quality, having a low solids content, to prevent excessive

wear. To ensure that the required water quality is achieved water

clarification plants (high rate thickeners) would be required.

The alternative method is to use single stage pumps capable of handling

dirty water. In comparison to multi-stage single-stage pumps have lower

head. It would therefore be necessary to have a larger number of pump

stations. To minimise the number of pump stations only pumps with high



head capacities are considered. In addition to this a number of pumps areconnected in series in each pump set.

The proposed dewatering method and layout is shown in Drawing D31112-

0001 in Appendix 1.

3.2.2 Pumping Volumes

Pumping volumes have been calculated based on the average flow

i t B d th lt f th l l ti th fl i

face pump at the lowest portion of the decline to the main pump or the next

development pump station in the decline. The development pipeline would

use 100 mm HDPE pipe.

A small pump situated in the development end, or other area where water

collects, will transfer the water to the skid mounted unit. These will be a

small electric or air operated pump. Refer to drawing D31112-0004 in



Appendix 1 for the schematic arrangement proposed.

An alternative arrangement that could be considered in place of the decline

piping system would be to install the main pipelines in boreholes. This would

have the effect of reducing power costs but would be more difficult to install

and may have a higher capital cost.

3.3 Flow Analysis

A d l f th i t d t i t ith l ti A b i

pressures developed from this scenario would typically result in the highest

pressures.

Following a pump set trip a pressure surge wave is generated resulting in

high and low pressure peaks. The pipeline gradient is low and pressures fall

below atmospheric pressure creating vapour cavities. The resulting collapse

of the cavities creates pressure spike superimposed on the surge pressure



wave. A vacuum breaker valve was modelled, being installed near to the topof the decline piping to assist in reducing the sub atmospheric pressure that

occur. This helped to reduce the spikes to acceptable values. Figure 3.1

shows the surge pressure profile for this event. The figure shows the typical

surge wave profile superimposed with the pressure spikes.

3500

3.4 Pumping Normal Operations

During normal pumping operations the average required flow rate is 27 l/s. The

main pumps will only be required to operate for a relatively short period each day.

On average each pump set will operate less than 7 hours each day. When

operating with a pump station dam having a live capacity of 250 m3 the pump set will

only need to start five times per day for these conditions.



3.5 Flood Pumping Operations

In order to match the pumping requirements of inflow from the high rainfall estimate

of 30,800 m3 in one day it will be necessary to have four pump sets operating. This

will not only include the pumps sets but will include four sets of piping, four sets of

electrical switchgear and cabling. Allowance must also be made for additional pump

chamber excavations. This arrangement would provide a high pumping capacity but

will be expensive.

O h i t ll th t t b di t d t t d Thi i t

3.6 Main Pump Station Pipeline

Steel piping is used for the main pump pipelines with flanged connections as this

type of piping is best suited for permanent installations where resistance to damage

is a requirement.

The main pumping system uses 200NB piping to match the flow requirement of



100 l/s per pump set. The operating pressures are:

Steady state: 2,345 kPa

Surge pressure: 3,000 kPa

Pipe wall thickness calculations have been made in accordance with the

requirements of the ANSI B31.3 piping code for process piping. The peak operating

pressure is the surge pressure of 3,000 kPa. The pipe wall thickness calculation

allows for a potential corrosion allowance of 2 mm. Pipe material is based on the

ti t ASTM A106 G d B ifi ti f ld d i i F thi

placed at the entrances to the dam and pump chamber to prevent inadvertent

access by mobile machinery.

4 CAPITAL COST ESTIMATE

Table 4.1 shows a summary of the capital cost estimate for the pumping system.



TABLE 4.1 – SUMMARY OF CAPITAL COST ESTIMATE

Description Capital (US$) Year Installed

Pump Station 1 $1 269 294 Year -2

Pump Station 2 $1 244 294 Year -1

Pump Station 3 $1 244 294 Year 2





TABLE 4.2 - BUDGET QUOTES FROM SUPPLIERS

Dewatering Pumps Weir Minerals Africa

Steel Piping Mining Pressure Systems

HDPE Piping Marley Pipes

Steelwork Steel Mech Engineering



Owing to the use of budget quotes for the major items at this level of study, a priceestimate accuracy of 1.05 was used.

The cost of the piping installation was made based on a provision of 25 percent of

the piping cost.

For this level of study, a project contingency of 20 percent and 12 percent for EPCMwas used.

TABLE 5.1 - OPERATING COST FACTORS APPLIED

DESCRIPTION FACTOR ITEMS

Electrical 0.25% Cabling, switchgear, transformers,

instrumentation and control

Fixed - Consumables 5% Not applicable

Fixed – High Wear 5% Pipes (HDPE and steel), fittings, pipe



supportsFixed – Low Wear 2% Storage dam, concrete items, manifolds,

pump chamber equipment

Rotating - High 20% All pumps

Rotating - Normal 10% Not applicable

The annual operating costs excludes labour.

TABLE 5.2 – SUMMARY OF OPERATING COSTS

Description Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Year 12 Year 13 Year 14 Year 15

Pump Station 1 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175 $73 175

Pump Station 2 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112

Pump Station 3 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112

Pump Station 4 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112

Pump Station 5 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112Pump Station 6 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112 $73 112

Pump Station 7 $73 112 $73 112 $73 112 $73 112 $73 112

Development Pumping $11 000 $13 750 $16 500 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250 $19 250

Emergency Flood Recovery $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962 $5 962

Contingency (20%) $18 027 $33 200 $33 750 $34 300 $48 922 $48 922 $63 545 $63 545 $78 167 $78 167 $92 790 $92 790 $107 412 $107 412 $107 412 $107 412 $107 412

Power Cos t $330 146 $612 538 $628 456 $644 374 $910 848 $910 848 $1 177 321 $1 177 321 $1 443 795 $1 443 795 $1 710 268 $1 710 268 $1 976 742 $1 976 742 $1 976 742 $1 976 742 $1 976 742

Total Operating Cost $438 311 $811 737 $830 955 $850 173 $1 204 382 $1 204 382 $1 558 590 $1 558 590 $1 912 798 $1 912 798 $2 267 006 $2 267 006 $2 621 214 $2 621 214 $2 621 214 $2 621 214 $2 621 214



© Royal HaskoningDHV (formerly Turgis Consulting Ltd) Page 12 of 12Report number 31112-02 11 Dec 2012



APPENDIX 1 – DRAWINGS









Item Description Quantity Unit Unit cost US$ Item TotalUS$

Annual MaintenanceUS$

Annual power CostUS$/kWh Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Year 12 Year 13 Year 14 Year 15

PUMP STATION1

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 114 125 125 2 503 - 127 628 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 2 594 38 905 778 - 39 683 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778

Concrete floor 50 m2 128 6 413 128 - 6 541 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 9 875 29 625 593 - 30 218 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593

Hi lift dirty waterpumps -C5Weir Envirotech- 3in series -2sets (1operating1

standby)

6 ea 40 024 240 143 48 029 265 197 553 368 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225

Installation of pump chamber equipment 30 shift 1 313 39 375 788 0 40 163 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 86 206 700 10 335 - 217 035 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 4 063 49 725 2 486 - 52 211 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 9 20 670 1 034 - 21 704 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 22 51 675 2 584 - 54 259 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 1 650 1 277 11 177 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927

Piping 100NB HDPE 1300 m 12 15 600 780 - 16 380 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 6 094 4 438 222 - 4 660 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 3 3 900 195 - 4 095 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 ea 150 000 150 000 375 - 150 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 218 750 218 750 547 - 219 297 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547

MV Cabling 400 m 150 60 000 150 - 60 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150

PUMP STATION2

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 114 125 125 2 503 - 127 628 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 2 594 38 905 778 - 39 683 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778

Concrete floor 50 m2 128 6 413 128 - 6 541 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 9 875 29 625 593 - 30 218 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593


standby)6 ea 40 024 240 143 48 029 265 197 553 368 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225

Installation of pump chamber equipment 30 shift 1 313 39 375 788 0 40 163 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 86 206 700 10 335 - 217 035 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 4 063 49 725 2 486 - 52 211 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 9 20 670 1 034 - 21 704 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 22 51 675 2 584 - 54 259 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 1 650 1 277 11 177 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927

Piping 100NB HDPE 1300 m 12 15 600 780 - 16 380 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 6 094 4 438 222 - 4 660 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 3 3 900 195 - 4 095 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 ea 150 000 150 000 375 - 150 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 218 750 218 750 547 - 219 297 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547

MV Cabling 233.3333333 m 150 35 000 88 - 35 088 88 88 88 88 88 88 88 88 88 88 88 88 88 88 88

PUMP STATION3

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 114 125 125 2 503 - 127 628 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 2 594 38 905 778 - 39 683 778 778 778 778 778 778 778 778 778 778 778 778

Concrete floor 50 m2 128 6 413 128 - 6 541 128 128 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 9 875 29 625 593 - 30 218 593 593 593 593 593 593 593 593 593 593 593 593


standby)6 ea 40 024 240 143 48 029 265 197 553 368 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225

Installation of pump chamber equipment 30 shift 1 313 39 375 788 0 40 163 788 788 788 788 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 86 206 700 10 335 - 217 035 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 4 063 49 725 2 486 - 52 211 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 9 20 670 1 034 - 21 704 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 22 51 675 2 584 - 54 259 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 1 650 1 277 11 177 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927

Piping 100NB HDPE 1300 m 12 15 600 780 - 16 380 780 780 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 6 094 4 438 222 - 4 660 222 222 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 3 3 900 195 - 4 095 195 195 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 ea 150 000 150 000 375 - 150 375 375 375 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 218 750 218 750 547 - 219 297 547 547 547 547 547 547 547 547 547 547 547 547

MV Cabling 233.3333333 m 150 35 000 88 - 35 088 88 88 88 88 88 88 88 88 88 88 88 88

PUMP STATION4

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 114 125 125 2 503 - 127 628 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 2 594 38 905 778 - 39 683 778 778 778 778 778 778 778 778 778 778

Concrete floor 50 m2 128 6 413 128 - 6 541 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 9 875 29 625 593 - 30 218 593 593 593 593 593 593 593 593 593 593


standby)6 ea 40 024 240 143 48 029 265 197 553 368 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225

Installation of pump chamber equipment 30 shift 1 313 39 375 788 - 40 163 788 788 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 86 206 700 10 335 - 217 035 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 4 063 49 725 2 486 - 52 211 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 9 20 670 1 034 - 21 704 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 22 51 675 2 584 - 54 259 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 1 650 1 277 11 177 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927

Piping 100NB HDPE 1300 m 12 15 600 780 - 16 380 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 6 094 4 438 222 - 4 660 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 3 3 900 195 - 4 095 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 ea 150 000 150 000 375 - 150 375 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 218 750 218 750 547 - 219 297 547 547 547 547 547 547 547 547 547 547

MV Cabling 233.3333333 m 150 35 000 88 - 35 088 88 88 88 88 88 88 88 88 88 88

PUMP STATION5

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 114 125 125 2 503 - 127 628 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 2 594 38 905 778 - 39 683 778 778 778 778 778 778 778 778

Concrete floor 50 m2 128 6 413 128 - 6 541 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 9 875 29 625 593 - 30 218 593 593 593 593 593 593 593 593


standby)6 ea 40 024 240 143 48 029 265 197 553 368 313 225 313 225 313 225 313 225 313 225 313 225 313 225 313 225

Installation of pump chamber equipment 30 shift 1 313 39 375 788 0 40 163 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 86 206 700 10 335 - 217 035 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 4 063 49 725 2 486 - 52 211 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 9 20 670 1 034 - 21 704 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 22 51 675 2 584 - 54 259 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 1 650 1 277 11 177 2 927 2 927 2 927 2 927 2 927 2 927 2 927 2 927

Piping 100NB HDPE 1300 m 12 15 600 780 - 16 380 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 6 094 4 438 222 - 4 660 222 222 222 222 222 222 222 222

Install Development piping 1300 m 3 3 900 195 - 4 095 195 195 195 195 195 195 195 195

Instrumentation and Control 1 ea 150 000 150 000 375 - 150 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 218 750 218 750 547 - 219 297 547 547 547 547 547 547 547 547

MV Cabling 233.3333333 m 150 35 000 88 - 35 088 88 88 88 88 88 88 88 88

PUMP STATION6

GUYANADEWATERING COST SUMMARY

ProjectCosting

SUMMARYOF EXPENDITUREOVERMINE LIFE



Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 1 14 125 125 2 503 - 127 628 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 2 594 38 905 778 - 39 683 778 778 778 778 778 778

Concrete floor 50 m2 128 6 413 128 - 6 541 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 9 875 29 625 593 - 30 218 593 593 593 593 593 593


standby)6 ea 40 024 240 143 48 029 265 197 553 368 313 225 313 225 313 225 313 225 313 225 313 225

Installation of pump chamber equipment 30 shift 1 313 39 375 788 0 40 163 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 86 206 700 10 335 - 217 035 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 4 063 49 725 2 486 - 52 211 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 9 20 670 1 034 - 21 704 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 22 51 675 2 584 - 54 259 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 1 650 1 277 11 177 2 927 2 927 2 927 2 927 2 927 2 927

Piping 100NB HDPE 1300 m 12 15 600 780 - 16 380 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 6 094 4 438 222 - 4 660 222 222 222 222 222 222

Install Development piping 1300 m 3 3 900 195 - 4 095 195 195 195 195 195 195

Instrumentation and Control 1 ea 150 000 150 000 375 - 150 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 218 750 218 750 547 - 219 297 547 547 547 547 547 547

MV Cabling 233.3333333 m 150 35 000 88 - 35 088 88 88 88 88 88 88

PUMP STATION7

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 1 14 125 125 2 503 - 127 628 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 2 594 38 905 778 - 39 683 778 778 778 778

Concrete floor 50 m2 128 6 413 128 - 6 541 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 9 875 29 625 593 - 30 218 593 593 593 593


standby)6 ea 40 024 240 143 48 029 265 197 553 368 313 225 313 225 313 225 313 225

Installation of pump chamber equipment 30 shift 1 313 39 375 788 0 40 163 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 86 206 700 10 335 - 217 035 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 4 063 49 725 2 486 - 52 211 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 9 20 670 1 034 - 21 704 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 22 51 675 2 584 - 54 259 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 1 650 1 277 11 177 2 927 2 927 2 927 2 927

Piping 100NB HDPE 1300 m 12 1 5 600 780 - 16 380 780 780 780 780

Piping support HDPE 0.7 tonnes 6 094 4 438 222 - 4 660 222 222 222 222

Install Development piping 1300 m 3 3 900 195 - 4 095 195 195 195 195

Instrumentation and Control 1 ea 150 000 150 000 375 - 150 375 375 375 375 375

Transformer and Switchgear 1 ea 218 750 218 750 547 - 219 297 547 547 547 547

MV Cabling 233.3333333 m 150 35 000 88 - 35 088 88 88 88 88

DevelopmentPumpingSkid mounted tank unit with VS pump 7 ea 13 750 96 250 19 250 111 427 170 923 93 341 112 009 130 677 130 677 130 677 130 677 130 677 130 677 130 677 130 677 130 677 130 677 130 677 130 677 130 677 130 677

0 - - - -

EmergencyFlood Recovery 0 - - - -

Flood recovery pump - Weir Envirotech - SHW 100-425 1500rpm 6 ea 2 594 15 562 3 112 - 18 675 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112

Piping 200NB HDPE 1200 m 48 57 000 2 850 - 59 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850

- - - -Project Contingency @ 20% 20% - 1 780 774 107 412 - 305 649 282 059 33750 34300 297781 48922 312404 63545 327026 78167 341648 92790 356271 107412 107412 107412 107412

Design & project control @ 12% 12% - 1 068 464 - - 1 068 464

Sub total 11 753 108 644 473 1 976 742 3 232 502 2304890 830955 850173 2697534 1204382 3051743 1558590 3405951 1912798 3760159 2267006 4114367 2621214 2621214 2621214 2621214

Item Description Quantity Unit Unit cost

US$

Item Total Capital

US$ Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Year 12 Year 13 Year 14 Year 15

PUMP STATION1

Waterstoragedam andpump stationexcavationandsupport 1100 m3 114 125 125 125 125

Concrete reinforced 15 m3 2 594 38 905 38 905

Concretefloor 50 m2 128 6 413 6 413

Supply of suction manifold piping - including valves 3.0 tonnes 9 875 29 625 29 625

Hi lift dirty waterpumps -C5Weir Envirotech- 3in series -2 sets (1operating1

standby)6 ea 40 024 240 143 240 143

Installation of pump chamber equipment 30 shift 1 313 39 375 39 375

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 86 206 700 206 700

Piping support steel 12.2 tonnes 4 063 49 725 49 725

Pipe fitings and valves 2400 m 9 20 670 20 670

Install 200 NB piping 2400 m 22 51 675 51 675

Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 8 250

Piping 100NB HDPE 1300 m 12 15 600 15 600

Piping support HDPE 0.7 tonnes 6 094 4 438 4 438

Install Development piping 1300 m 3 3 900 3 900

Instrumentation and Control 1 ea 150 000 150 000 150 000

Transformer and Switchgear 1 ea 218 750 218 750 218 750

MV Cabling 400 m 150 60 000 60 000

PUMP STATION2

Waterstoragedam 250m3 andpump stationexcavationandsupport 1100 m3 114 125 125 125 125





standby)6 ea 40 024 240 143 240 143







Piping 100NB HDPE 1300 m 12 15 600 15 600



Instrumentation and Control 1 m 150 000 150 000 150 000

Transformer and Switchgear 1 m 218 750 218 750 218 750

MV Cabling 233.3333333 m 150 35 000 35 000

PUMP STATION3






standby)6 ea 40 024 240 143 240 143






Vertical spindle pumps - sumps and piping 1 ea 8 250 8 250 8 250Piping 100NB HDPE 1300 m 12 1 5 600 15 600





MV Cabling 233.3333333 m 150 35 000 35 000

PUMP STATION4






standby)6 ea 40 024 240 143 240 143







Piping 100NB HDPE 1300 m 12 15 600 15 600





MV Cabling 233.3333333 m 150 35 000 35 000

PUMP STATION5






standby)6 ea 40 024 240 143 240 143







Piping 100NB HDPE 1300 m 12 15 600 15 600





MV Cabling 233.3333333 m 150 35 000 35 000

PUMP STATION6

Waterstoragedam 250m

3

andpump stationexcavationandsupport 1100 m

3

114 125 125 125 125Concrete reinforced 15 m3 2 594 38 905 38 905

CAPITAL EXPENDITUREOVER MINELIFE

ProjectCosting







standby)6 ea 40 024 240 143 240 143







Piping 100NB HDPE 1300 m 12 15 600 15 600





MV Cabling 233.3333333 m 150 35 000 35 000

PUMP STATION7






standby)6 ea 40 024 240 143 240 143



Piping support steel 12.2 t onnes 4 063 49 725 49 725




Piping 100NB HDPE 1300 m 12 15 600 15 600





MV Cabling 233.3333333 m 150 35 000 35 000

DevelopmentPumping

Skid mounted tank unit with VS pump 7 ea 13 750 96 250 96 250

EmergencyFlood Recovery

Flood recovery pump -Weir Envirotech -SHW 100-425 1500rpm 6 ea 2 594 15 562 15 562

Piping 200NB HDPE 1200 m 48 57 000 57 000

Project Contingency @ 20% 20% - 1 780 774 287 621 248 859 - - 248859 - 248859 - 248859 - 248859 - 248859 - - - -

Design & project control @ 12% 12% - 1 068 464 1 068 464

Sub total 11 753 108 2 794 192 1493153 - - 1493153 - 1493153 - 1493153 - 1493153 - 1493153 - - - -

Item Description Quantity Unit Annual Maintenance

US$ Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Year 12 Year 13 Year 14 Year 15

PUMP STATION1

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778

Concretefloor 50 m2 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593Hi lift dirty waterpumps -C5WeirEnvirotech-3 inseries -2 sets (1operating1

standby)6 ea 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029

Installation of pump chamber equipment 30 shift 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650

Piping 100NB HDPE 1300 m 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 ea 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547

MV Cabling 400 m 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150 150

PUMP STATION2

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778 778

Concretefloor 50 m2 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593 593

Hi lift dirty waterpumps -C5WeirEnvirotech-3 inseries -2 sets (1operating1

standby)6 ea 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029

Installation of pump chamber equipment 30 shift 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584Vertical spindle pumps - sumps and piping 1 ea 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650

Piping 100NB HDPE 1300 m 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 m 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 m 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547 547

MV Cabling 233.3333333 m 88 88 88 88 88 88 88 88 88 88 88 88 88 88 88 88 88

-

PUMP STATION3 -

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 778 778 778 778 778 778 778 778 778 778 778 778 778 778

Concretefloor 50 m2 128 128 128 128 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 593 593 593 593 593 593 593 593 593 593 593 593 593 593


standby)6 ea 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029

Installation of pump chamber equipment 30 shift 788 788 788 788 788 788 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650

Piping 100NB HDPE 1300 m 780 780 780 780 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 222 222 222 222 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 195 195 195 195 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 m 375 375 375 375 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 m 547 547 547 547 547 547 547 547 547 547 547 547 547 547MV Cabling 233.3333333 m 88 88 88 88 88 88 88 88 88 88 88 88 88 88

-PUMP STATION4 -

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 778 778 778 778 778 778 778 778 778 778 778 778

Concretefloor 50 m2 128 128 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 593 593 593 593 593 593 593 593 593 593 593 593


standby)6 ea 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029

Installation of pump chamber equipment 30 shift 788 788 788 788 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650

Piping 100NB HDPE 1300 m 780 780 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 222 222 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 195 195 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 ea 375 375 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 547 547 547 547 547 547 547 547 547 547 547 547

MV Cabling 233.3333333 m 88 88 88 88 88 88 88 88 88 88 88 88

PUMP STATION5

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 778 778 778 778 778 778 778 778 778 778

Concretefloor 50 m2 128 128 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 593 593 593 593 593 593 593 593 593 593


standby)6 ea 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029

Installation of pump chamber equipment 30 shift 788 788 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486Pipe fitings and valves 2400 m 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650

Piping 100NB HDPE 1300 m 780 780 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 222 222 222 222 222 222 222 222 222 222

Install Development piping 1300 m 195 195 195 195 195 195 195 195 195 195

Instrumentation and Control 1 ea 375 375 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 547 547 547 547 547 547 547 547 547 547

MV Cabling 233.3333333 m 88 88 88 88 88 88 88 88 88 88

PUMP STATION6

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 2 503 2 503 2 503 2 503 2 503 2 503 2 503 2 503

Concrete reinforced 15 m3 778 778 778 778 778 778 778 778

Concretefloor 50 m2 128 128 128 128 128 128 128 128

Supply of suction manifold piping - including valves 3.0 tonnes 593 593 593 593 593 593 593 593


standby)6 ea 48 029 48 029 48 029 48 029 48 029 48 029 48 029 48 029

Installation of pump chamber equipment 30 shift 788 788 788 788 788 788 788 788

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 10 335 10 335 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 2 486 2 486 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 1 034 1 034 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 2 584 2 584 2 584 2 584 2 584 2 584 2 584 2 584Vertical spindle pumps - sumps and piping 1 ea 1 650 1 650 1 650 1 650 1 650 1 650 1 650 1 650

Pi i 100NB HDPE 1300 780 780 780 780 780 780 780 780


ProjectCosting

OPERATING EXPENDITUREOVERMINE LIFE(excl. Power and Labour)



Piping 100NB HDPE 1300 m 780 780 780 780 780 780 780 780

Piping support HDPE 0.7 tonnes 222 222 222 222 222 222 222 222

Install Development piping 1300 m 195 195 195 195 195 195 195 195Instrumentation and Control 1 ea 375 375 375 375 375 375 375 375

Transformer and Switchgear 1 ea 547 547 547 547 547 547 547 547

MV Cabling 233.3333333 m 88 88 88 88 88 88 88 88

PUMP STATION7

Waterstoragedam 250m3 andpump stationexcavationand support 1100 m3 2 503 2 503 2503 2503 2503 2503


Concretefloor 50 m2 128 128 128 128 128 128



standby)6 ea 48 029 48 029 4 8 029 4 8 029 4 8 029 4 8 029


Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m 10 335 10 335 10 335 10 335 10 335 10 335

Piping support steel 12.2 tonnes 2 486 2 486 2 486 2 486 2 486 2 486

Pipe fitings and valves 2400 m 1 034 1 034 1 034 1 034 1 034 1 034

Install 200 NB piping 2400 m 2 584 2 584 2 584 2 584 2 584 2 584

Vertical spindle pumps - sumps and piping 1 ea 1 650 1 650 1 650 1 650 1 650 1 650

Piping 100NB HDPE 1300 m 780 780 780 780 780 780

Piping support HDPE 0.7 t onnes 222 222 222 222 222 222

Install Development piping 1300 m 195 195 195 195 195 195

Instrumentation and Control 1 ea 375 375 375 375 375 375

Transformer and Switchgear 1 ea 547 547 547 547 547 547

MV Cabling 233.3333333 m 88 88 88 88 88 88

DevelopmentPumping

Skid mounted tank unit with VS pump 7 ea 19 250 11 000 13 750 16 500 19 250 19 250 19 250 19 250 19 250 19 250 19 250 19 250 19 250 19 250 19 250 19 250 19 250 19 2500 -

EmergencyFlood Recovery 0 -

Flood recovery pump - Weir Envirotech - SHW 100-425 1500rpm 6 ea 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112 3 112

Piping 200NB HDPE 1200 m 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850 2 850

-

Project Contingency @ 20% 20% 107 412 18 027 33 200 33750 34300 48922 48922 63545 63545 78167 78167 92790 92790 107412 107412 107412 107412 107412

Sub total 644 473 108 165 199199 202499 205799 293534 293534 381269 381269 469003 469003 556738 556738 644473 644473 644473 644473 644473

Item Description Quantity Unit Annual power Cost Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Year 12 Year 13 Year 14 Year 15

PUMP STATION1

Waterstoragedam 250m3 andpump stationexcavationandsupport 1100 m3 - - - - - - - - - - - - - - - - - -

Concrete reinforced 15 m3 - - - - - - - - - - - - - - - - - -

Concretefloor 50 m2 - - - - - - - - - - - - - - - - - -

Supply of suction manifold piping - including valves 3.0 tonnes - - - - - - - - - - - - - - - - - -

Hi lift dirty waterpumps -C5WeirEnvirotech-3 inseries -2 sets (1operating1standby)

6 ea 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197

Installation of pump chamber equipment 30 shift 0 - - - - - - - - - - - - - - - - -

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m - - - - - - - - - - - - - - - - - -

Piping support steel 12.2 t onnes - - - - - - - - - - - - - - - - - -

Pipe fitings and valves 2400 m - - - - - - - - - - - - - - - - - -

Install 200 NB piping 2400 m - - - - - - - - - - - - - - - - - -

Vertical spindle pumps - sumps and piping 1 ea 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277

Piping 100NB HDPE 1300 m - - - - - - - - - - - - - - - - - -

Piping support HDPE 0.7 t onnes - - - - - - - - - - - - - - - - - -

Install Development piping 1300 m - - - - - - - - - - - - - - - - - -

Instrumentation and Control 1 ea - - - - - - - - - - - - - - - - - -

Transformer and Switchgear 1 ea - - - - - - - - - - - - - - - - - -

MV Cabling 400 m - - - - - - - - - - - - - - - - - -

PUMP STATION2

Waterstoragedam 250m3 andpump stationexcavationandsupport 1100 m3 - - - - - - - - - - - - - - - - -

Concrete reinforced 15 m3 - - - - - - - - - - - - - - - - -

Concretefloor 50 m2 - - - - - - - - - - - - - - - - -

Supply of suction manifold piping - including valves 3.0 tonnes - - - - - - - - - - - - - - - - -


standby)6 ea 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197

Installation of pump chamber equipment 30 shift 0 - - - - - - - - - - - - - - - -

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m - - - - - - - - - - - - - - - - -

Piping support steel 12.2 tonnes - - - - - - - - - - - - - - - - -

Pipe fitings and valves 2400 m - - - - - - - - - - - - - - - - -

Install 200 NB piping 2400 m - - - - - - - - - - - - - - - - -

Vertical spindle pumps - sumps and piping 1 ea 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277

Piping 100NB HDPE 1300 m - - - - - - - - - - - - - - - - -

Piping support HDPE 0.7 t onnes - - - - - - - - - - - - - - - - -

Install Development piping 1300 m - - - - - - - - - - - - - - - - -

Instrumentation and Control 1 ea - - - - - - - - - - - - - - - - - -

Transformer and Switchgear 1 ea - - - - - - - - - - - - - - - - - -

MV Cabling 233.3333333 m - - - - - - - - - - - - - - - - - -

0 -

PUMP STATION3 0 -

Waterstoragedam 250m3 andpump stationexcavationandsupport 1100 m3 - - - - - - - - - - - - - -

Concrete reinforced 15 m3 - - - - - - - - - - - - - -

Concretefloor 50 m2 - - - - - - - - - - - - - -

Supply of suction manifold piping - including valves 3.0 tonnes - - - - - - - - - - - - - -


standby)6 ea 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197

Installation of pump chamber equipment 30 shift 0 - - - - - - - - - - - - -

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m - - - - - - - - - - - - - -

Piping support steel 12.2 tonnes - - - - - - - - - - - - - -

Pipe fitings and valves 2400 m - - - - - - - - - - - - - -

Install 200 NB piping 2400 m - - - - - - - - - - - - - -

Vertical spindle pumps - sumps and piping 1 ea 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277

Piping 100NB HDPE 1300 m - - - - - - - - - - - - - -

Piping support HDPE 0.7 tonnes - - - - - - - - - - - - - -Install Development piping 1300 m - - - - - - - - - - - - - -

Instrumentation and Control 1 ea - - - - - - - - - - - - - -

Transformer and Switchgear 1 ea - - - - - - - - - - - - - -

MV Cabling 233.3333333 m - - - - - - - - - - - - - -

0 -

PUMP STATION4 0 -

Waterstoragedam 250m3 andpump stationexcavationandsupport 1100 m3 - - - - - - - - - - - -

Concrete reinforced 15 m3 - - - - - - - - - - - -

Concretefloor 50 m2 - - - - - - - - - - - -

Supply of suction manifold piping - including valves 3.0 tonnes - - - - - - - - - - - -


standby)6 ea 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197

Installation of pump chamber equipment 30 shift - - - - - - - - - - - -

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m - - - - - - - - - - - -

Piping support steel 12.2 tonnes - - - - - - - - - - - -

Pipe fitings and valves 2400 m - - - - - - - - - - - -

Install 200 NB piping 2400 m - - - - - - - - - - - -

Vertical spindle pumps - sumps and piping 1 ea 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277

Piping 100NB HDPE 1300 m - - - - - - - - - - - -

Piping support HDPE 0.7 tonnes - - - - - - - - - - - -

Install Development piping 1300 m - - - - - - - - - - - -

Instrumentation and Control 1 ea - - - - - - - - - - - -

Transformer and Switchgear 1 ea - - - - - - - - - - - -

MV Cabling 233.3333333 m - - - - - - - - - - - -

0 -

PUMP STATION5 0 -

Waterstoragedam 250m3 andpump stationexcavationandsupport 1100 m3 - - - - - - - - - -

Concrete reinforced 15 m3 - - - - - - - - - -

Concretefloor 50 m2 - - - - - - - - - -

Supply of suction manifold piping - including valves 3.0 tonnes - - - - - - - - - -


standby)6 ea 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197

Installation of pump chamber equipment 30 shift 0 - - - - - - - - -

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m - - - - - - - - - -

Piping support steel 12.2 tonnes - - - - - - - - - -

Pipe fitings and valves 2400 m - - - - - - - - - -

Install 200 NB piping 2400 m - - - - - - - - - -

Vertical spindle pumps - sumps and piping 1 ea 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277

Piping 100NB HDPE 1300 m - - - - - - - - - -

Piping support HDPE 0.7 tonnes - - - - - - - - - -

Install Development piping 1300 m - - - - - - - - - -

Instrumentation and Control 1 ea - - - - - - - - - -

Transformer and Switchgear 1 ea - - - - - - - - - -

MV Cabling 233.3333333 m - - - - - - - - - -

0 -

PUMP STATION6 0 -

Waterstoragedam 250m3 andpump stationexcavationandsupport 1100 m3 - - - - - - - -

Concrete reinforced 15 m3 - - - - - - - -

Concretefloor 50 m2 - - - - - - - -Supply of suction manifold piping - including valves 3.0 tonnes - - - - - - - -


ProjectCosting

POWEREXPENDITUREOVERMINELIFE




standby)6 ea 265 197 265 197 265 197 265 197 265 197 265 197 265 197 265 197

Installation of pump chamber equipment 30 shift 0 - - - - - - -

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m - - - - - - - -

Piping support steel 12.2 tonnes - - - - - - - -

Pipe fitings and valves 2400 m - - - - - - - -

Install 200 NB piping 2400 m - - - - - - - -

Vertical spindle pumps - sumps and piping 1 ea 1 277 1 277 1 277 1 277 1 277 1 277 1 277 1 277

Piping 100NB HDPE 1300 m - - - - - - - -

Piping support HDPE 0.7 tonnes - - - - - - - -

Install Development piping 1300 m - - - - - - - -

Instrumentation and Control 1 ea - - - - - - - -

Transformer and Switchgear 1 ea - - - - - - - -

MV Cabling 233.3333333 m - - - - - - - -

0 -

PUMP STATION7 0 -

Waterstoragedam 250m3 andpump stationexcavationandsupport 1100 m3 - - - - - -

Concrete reinforced 15 m3 - - - - - -

Concretefloor 50 m2 - - - - - -

Supply of suction manifold piping - including valves 3.0 tonnes - - - - - -


standby)6 ea 265 197 265 197 265 197 265 197 265 197 265 197

Installation of pump chamber equipment 30 shift 0 - - - - -

Supply of 200 NB 6mm wt class 2500 pump delivery piping in declines 2400 m - - - - - -

Piping support steel 12.2 tonnes - - - - - -

Pipe fitings and valves 2400 m - - - - - -

Install 200 NB piping 2400 m - - - - - -

Vertical spindle pumps - sumps and piping 1 ea 1 277 1 277 1 277 1 277 1 277 1 277

Piping 100NB HDPE 1300 m - - - - - -

Piping support HDPE 0.7 tonnes - - - - - -

Install Development piping 1300 m - - - - - -

Instrumentation and Control 1 e a - - - - - -

Transformer and Switchgear 1 ea - - - - - -

MV Cabling 233.3333333 m - - - - - -

0 -

DevelopmentPumping 0 -

Skid mounted tank unit with VS pump 7 ea 111 427 63 673 79 591 95 509 111 427 111 427 111 427 111 427 111 427 111 427 111 427 111 427 111 427 111 427 111 427 111 427 111 427 111 427

0 -

EmergencyFlood Recovery 0 -

Flood recovery pump - Weir Envirotech - SHW 100-425 1500rpm 6 ea - - - - - - - - - - - - - - - - - -

Piping 200NB HDPE 1200 m - - - - - - - - - - - - - - - - - -

-

Project Contingency @ 20% n/a -

Sub total US$ 330146 612538 628456 644374 910848 910848 1177321 1177321 1443795 1443795 1710268 1710268 1976742 1976742 1976742 1976742 1976742


Documents

NI 43-101 Technical Report Updated Feasibility Study Aurora Gold Project