SCOPING STUDY - Star Gold€¦ · This Scoping Study has identified potentially mineable resources...

SCOPING STUDY Longstreet Gold Project, Nye County, Nevada USA

Prepared for Star Gold Inc.

May 27, 2014

A-Z Mining Professionals Limited

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Star Gold Inc. Longstreet Project Scoping Study i

Table of Contents

1 EXECUTIVE SUMMARY ........................................................................................................................... 1

1.1 Conclusions .................................................................................................................................... 6

1.2 Recommendations ........................................................................................................................ 6

2 INTRODUCTION ..................................................................................................................................... 8

3 LOCATION & HISTORY ........................................................................................................................... 9

4 GEOLOGY ............................................................................................................................................. 14

4.1 Regional Geology ......................................................................................................................... 14

4.2 Property Geology......................................................................................................................... 14

5 RESOURCES .......................................................................................................................................... 15

5.1 Resource Comparison.................................................................................................................. 15

5.2 Block Model Review .................................................................................................................... 15

5.2.1 Database .............................................................................................................................. 16

5.3 Mineral Wire Framing ................................................................................................................. 17

5.4 Sample Capture and Compositing ............................................................................................... 18

5.5 Variography and Variogram Modelling ....................................................................................... 18

5.6 Block Model Definition ................................................................................................................ 19

5.7 Estimation Search Criteria ........................................................................................................... 19

5.8 Statistical and Visual Checks of Model ........................................................................................ 20

5.9 Resource Reporting ..................................................................................................................... 21

5.10 Conclusion ................................................................................................................................... 22

5.11 Geology and resource estimate recommendations .................................................................... 22

6 GEOTECHNICAL AND HYDROGEOLOGY ............................................................................................... 23

6.1 Regional Hydrogeology ............................................................................................................... 23

6.2 Local Hydrogeology ..................................................................................................................... 23

6.3 Mine Site Hydrogeology .............................................................................................................. 25

6.4 Hydrogeology conclusion and recommendation ........................................................................ 26

6.5 Geotechnical ................................................................................................................................ 27

7 MINING ................................................................................................................................................ 28

7.1 Potentially Mineable Mineral Resource – Pit Optimization ........................................................ 28

7.2 Mining Method ............................................................................................................................ 35

7.3 Mining Schedule .......................................................................................................................... 35

7.4 Mine Closure ............................................................................................................................... 36

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Star Gold Inc. Longstreet Project Scoping Study ii

8 METALLURGICAL TESTWORK and MINERAL PROCESSING .................................................................. 38

8.1 Historical metallurgical sampling and test work ......................................................................... 38

8.2 2013 Sampling for metallurgical testwork .................................................................................. 39

8.3 Gold-Silver Mineralogy ................................................................................................................ 43

8.3.1 Sulfide Mineralogy ............................................................................................................... 43

8.3.2 Oxide Mineralogy ................................................................................................................ 43

8.4 2013 Metallurgical Test Program ................................................................................................ 43

8.4.1 Section Sample Assays ......................................................................................................... 44

8.4.2 Bottle Roll Testing ............................................................................................................... 44

8.4.3 Column Leach Testing .......................................................................................................... 46

8.4.4 Comminution Tests ............................................................................................................. 49

8.4.5 Mineral Processing Conclusions: ......................................................................................... 50

8.4.6 Mineral Processing Recommendations ............................................................................... 50

8.5 Process Engineering and Design .................................................................................................. 51

8.6 Process Description ..................................................................................................................... 55

The process of winning gold from the mined ore involves a number of individual activities,

described in detail below. ................................................................................................................... 55

8.6.1 Crushing ............................................................................................................................... 55

8.6.2 Double Deck Screen ............................................................................................................. 55

8.6.3 Lime Addition ...................................................................................................................... 55

8.6.4 Heap Leach Pad Stacking ..................................................................................................... 55

8.8.3 Leachate Distribution & Collection...................................................................................... 56

8.6.5 Solution Ponds ..................................................................................................................... 57

8.7 Adsorption, Desorption and Refining (ADR) Facility ................................................................... 57

8.7.1 Adsorption circuit ................................................................................................................ 57

8.7.2 Carbon acid washing ........................................................................................................... 58

8.7.3 Desorption circuit ................................................................................................................ 58

8.7.4 Carbon thermal regeneration.............................................................................................. 59

8.7.5 Refining ................................................................................................................................ 59

8.7.6 Water services ..................................................................................................................... 59

8.7.7 Reagents .............................................................................................................................. 60

8.7.8 Assay laboratory .................................................................................................................. 60

8.8 ADR plant manpower .................................................................................................................. 60

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Star Gold Inc. Longstreet Project Scoping Study iii

9 INFRASTRUCTURE ................................................................................................................................ 61

9.1 Site Access ................................................................................................................................... 61

9.2 Power and Power Distribution .................................................................................................... 61

9.3 Site roads ..................................................................................................................................... 61

9.4 Surface support buildings ............................................................................................................ 61

9.5 Other services .............................................................................................................................. 62

9.6 Area support services .................................................................................................................. 62

9.7 General and administrative ......................................................................................................... 62

9.7.1 Administration ..................................................................................................................... 62

9.7.2 Procurement ........................................................................................................................ 63

9.7.3 Human Resources ................................................................................................................ 63

9.7.4 Security ................................................................................................................................ 63

9.7.5 Manpower ........................................................................................................................... 64

10 HYDROLOGY .................................................................................................................................... 65

10.1 Water Sources ............................................................................................................................. 65

10.2 Water Usage ................................................................................................................................ 66

10.3 Dewatering .................................................................................................................................. 67

11 ENVIRONMENTAL AND PERMITTING .............................................................................................. 68

11.1 Permitting Process ....................................................................................................................... 68

11.1.1 US Forest Service ................................................................................................................. 68

11.1.2 Bureau of Land Management .............................................................................................. 69

11.1.3 Nevada Division of Environmental Protection (and other Agencies as noted) ................... 69

11.1.4 Other Permits ...................................................................................................................... 70

11.2 Timing of Approvals ..................................................................................................................... 70

11.3 Inventoried Roadless Area .......................................................................................................... 70

11.4 Greater Sage-grouse .................................................................................................................... 71

11.5 Cultural resources ....................................................................................................................... 74

11.6 Environmental and permitting conclusions ................................................................................ 74

12 PROJECT DEVELOPMENT SCHEDULE ............................................................................................... 75

13 CAPITAL EXPENDITURES ESTIMATES ............................................................................................... 77

13.1 Basis for estimates ...................................................................................................................... 77

13.2 Mining .......................................................................................................................................... 77

13.3 Heap leach and processing plant................................................................................................. 77

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Star Gold Inc. Longstreet Project Scoping Study iv

13.4 Infrastructure and support facilities ............................................................................................ 78

13.5 Owners costs ............................................................................................................................... 79

13.6 Total capital expenditures ........................................................................................................... 79

13.7 Sustaining capital ......................................................................................................................... 80

13.8 Closure costs ................................................................................................................................ 80

14 OPERATING COST ESTIMATES ......................................................................................................... 81

14.1 Basis for estimates ...................................................................................................................... 81

14.2 Mining .......................................................................................................................................... 81

14.3 Heap leach and gold recovery plant ............................................................................................ 81

14.4 General & Administration Costs .................................................................................................. 83

14.5 Dore transport and refining charges ........................................................................................... 85

14.6 Project total operating costs ....................................................................................................... 85

14.7 Exclusions .................................................................................................................................... 86

15 ECONOMIC ANALYSIS ...................................................................................................................... 87

15.1 Basis for analysis .......................................................................................................................... 87

15.2 Metal price derivation ................................................................................................................. 88

15.3 Financial returns .......................................................................................................................... 88

15.4 Sensitivity analysis ....................................................................................................................... 91

15.5 Economic interpretations and conclusions ................................................................................. 94

16 PROJECT RISK ASSESSMENT ............................................................................................................ 95

17 CONCLUSIONS AND RECOMMENDATIONS ..................................................................................... 96

17.1 Conclusions .................................................................................................................................. 96

17.2 Recommendations ...................................................................................................................... 97

18 REFERENCES .................................................................................................................................... 99

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Star Gold Inc. Longstreet Project Scoping Study v

LIST OF TABLES

Table 1.1 Longstreet Project Study Parameters ............................................................................................ 2

Table 1.2 Pre-production Capital Costs ......................................................................................................... 2

Table 1.3 Summary of Operating Costs ......................................................................................................... 3

Table 1.4 Longstreet Project Returns ............................................................................................................ 3

Table 5.1 Mineral Resources - Longstreet Gold Project .............................................................................. 15

Table 5.2 Comparison of Agnerian and Noland Block Models .................................................................... 16

Table 7.1 Floating cone pit optimization parameters ................................................................................. 28

Table 7.2 In-pit Mineral Resource Estimate ................................................................................................ 29

Table 7.3 Mine Schedule ............................................................................................................................. 36

Table 8.1 Metallurgical Testwork Results, (c. 1988) .................................................................................... 38

Table 8.2 Gold Head Assays and Head Grade Comparisons ....................................................................... 44

Table 8.3 Bottle Roll Test Results, 2013 ...................................................................................................... 45

Table 8.4 Summary Metallurgical Test Results ........................................................................................... 47

Table 8.5 Process Design Criteria ................................................................................................................ 51

Table 9.1 G&A Personnel Complement ....................................................................................................... 64

Table 13.1 ADR Plant Capital Expenditure Estimate ................................................................................... 78

Table 14.1 Processing Plant Operating Cost ............................................................................................... 82

Table 14.2 General and Administrative Operating Cost Components ........................................................ 84

Table 14.3 G&A Manpower Costs ............................................................................................................... 85

Table 14.4 Project Operating Cost Summary .............................................................................................. 86

Table 15.1 Cash Flow Model Inputs ............................................................................................................ 87

Table 15.2 Longstreet Gold Project Cash Flow Model ................................................................................ 90

Table 15.3 Cash Flow Summary................................................................................................................... 91

Table 15.4 Pre-Tax Sensitivity Analysis ....................................................................................................... 91

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Star Gold Inc. Longstreet Project Scoping Study vi

LIST OF FIGURES

Figure 1.1 Pre-tax IRR Sensitivities ................................................................................................................ 4

Figure 1.2 Project NPV Sensitivities ............................................................................................................... 4

Figure 1.3 Project IRR Sensitivity to Gold Price ............................................................................................. 5

Figure 1.4 Project NPV Sensitivity to Gold Price ........................................................................................... 5

Figure 3.1 Longstreet Gold Project Location ............................................................................................... 10

Figure 3.2 Longstreet Gold Project Satellite View (north to top of page) ................................................... 11

Figure 3.3 Longstreet Gold Project - Main Zone ......................................................................................... 12

Figure 3.4 Cross-section through Upper Adit .............................................................................................. 13

Figure 5.1 Sample section of Agnerian geological model ........................................................................... 17

Figure 5.2 Sample section through Noland model (same section) ............................................................. 18

Figure 5.3 Detailed section through Noland geologic model ...................................................................... 21

Figure 6.1 Great Basin Hydrogeology .......................................................................................................... 24

Figure 6.2 Water well search location recommendations .......................................................................... 26

Figure 7.1 Pit shell aerial view ..................................................................................................................... 30

Figure 7.2 Pit shell longitudinal section, 13,935,430E ................................................................................ 31

Figure 7.3 Pit shell longitudinal section 13,935,430E with topography ...................................................... 32

Figure 7.4 Pit shell cross-section 13,934,000N............................................................................................ 33

Figure 7.5 Pit shell cross-section 13,935,145N............................................................................................ 34

Figure 8.1 Location of Underground Adit Samples ..................................................................................... 41

Figure 8.2 Approximate Location of the Three Surface Sampling Pits (shown in red)................................ 42

Figure 8.3 Crush Size Versus Metal Recovery............................................................................................. 46

Figure 8.4 Surface Master composite leach kinetics .................................................................................. 48

Figure 8.5 Underground master composite leach kinetics ......................................................................... 48

Figure 8.6 Master blend composite leach kinetics ...................................................................................... 49

Figure 8.7 Flowsheet block diagram ............................................................................................................ 53

Figure 8.8 Flow sheet diagram .................................................................................................................... 54

Figure 11.1 Greater Sage-grouse habitat .................................................................................................... 73

Figure 12.1 Longstreet Gold Project engineering and development schedule ........................................... 76

Figure 15.1 Gold price trend 2006 -2014 .................................................................................................... 88

Figure 15.2 Pre-Tax IRR Sensitivities ........................................................................................................... 92

Figure 15.3 NPV10 Sensitivity Analysis ......................................................................................................... 92

Figure 15.4 Project IRR Sensitivity to Gold Price ......................................................................................... 93

Figure 15.5 Project NPV Sensitivity to Gold Price ....................................................................................... 93

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Star Gold Inc. Longstreet Project Scoping Study 1

1 EXECUTIVE SUMMARY

The mineral resources for the Longstreet Project were estimated by Agnerian Consulting Ltd.

(and reported in the NI 43-101 report entitled “Technical Report on the Longstreet Gold-Silver

Property, Nevada”, dated December 15, 2013). The geological block model developed by

Agnerian was used by A-Z Mining Professionals Ltd. (AMPL) to form the basis of the scoping

study reported herein.

This report has been prepared in Imperial units of measure. Key conversions are:

1 metric tonne = 1.1025 short tons 1 ounce Troy = 31.1035 grams 1 ounce Troy/ton = 34.3 grams/tonne

This Scoping Study has identified potentially mineable resources of 4.4 million short tons at

0.022 ounces gold per short ton and 0.53 ounces silver per short ton. The study includes

Indicated and Inferred mineral resources that would be contained in a designed open pit shell,

as allowed for studies of this nature by Canadian securities regulators. The reader is cautioned

that US SEC Industry Guide 7 disallows the economic studies of mineral properties on

mineralized material of lesser classification than Proven and Probable Reserves.

The deposit would be mined by open pit with the gold and silver extracted by heap leach and a

gold recovery plant. Infrastructure facilities would be minimized but include a small surface

shop, warehouse, office complex and water treatment facility.

The mine would operate at 1 million tons per annum and produce approximately 82,450 ounces

of gold and 348,200 ounces silver over its operating life. Based upon metallurgical testwork

conducted in 2013, gold recovery is expected to be 86% and silver recovery 15%. Recovered

(payable) silver represents only 7% of the total revenue of the mine. The parameters used in the

cashflow model are shown in Table 1.1.

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Table 1.1 Longstreet Project Study Parameters

Component Parameter

Potentially Mineable Resource (Indicated & Inferred), including mining dilution & recovery

4.4 million tons

Estimated Mining Dilution 5 percent @ 0% grade

Average mill head grade, gold 0.022 opt

Average mill head grade, silver 0.53 opt

Payable gold 82,450 ounces

Payable silver 348,200 ounces

Average long-term gold price $1,350 per ounce

Average long term silver price $24.00 per ounce

Pre-Production Capital, including Working Capital $25.4 million

Total Sustaining Capital $0

Closure Cost $1 million

Royalty 3% NSR

Estimated Operating Costs ($/ton) $14.87

Life of Mine 4.4 Years

Metal prices were provided by Star Gold. The gold price used in the study is lower than the 3-

year trailing average price, a common long-term price indicator, and within 5% of the current

price.

A summary of estimated capital costs is presented in Table 1.2

Table 1.2 Pre-production Capital Costs

Cost Component Expenditure

($US)

Permitting $ 2,000,000

Mine $ 220,000

Heap Leach Pad Processing Plant

$ 2,250,000 $ 8,097,000

Surface Infrastructure & Mobile Equipment $ 4,000,000

EPCM, Contractor O/H & Owners Costs Contingency

$ 2,535,000 $ 2,565,000

Total Capital Expenditures $21,667,000

Working Capital $ 3,690,000

TOTAL EXPENDITURES $25,357,000

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The heap leach pad and gold recovery plant represent 40% of the estimated total pre-

production capital expenditure. The estimate includes $4.2 million in contingency and working

capital allocations.

The estimated operating cost for mining, ore processing and general and administrative costs

are itemized in Table 1.3, expressed in US dollars per short ton processed. The operation is

expected to incur a cash cost of US$808 per ounce, exclusive of silver credits, or $760 per ounce

net of silver credits.

Table 1.3 Summary of Operating Costs

Department Cost

($US/ton)

Mine $ 9.09

Processing & Environmental $ 3.65

Surface Dept. and G&A $ 2.13

TOTAL $14.87

A 3% NSR royalty is held on the property by MinQuest, the vendor to Star Gold, and has been

factored into the cash flow model.

Economic analysis has indicated a robust return as shown in Table 1.4, with an estimated IRR of

29% and a present value of $13.3 million at a 10% discount factor, on a pre-tax basis.

Table 1.4 Longstreet Project Returns

Metric Value

Undiscounted Net Revenue $119 million

Undiscounted Cashflow $ 29 million

NPV (10%) $ 13 million


IRR 29%

Payback Period 2.7 years

The IRR and NPV sensitivities to variations in key parameters are depicted graphically in Figures

1.1 and 1.2. The IRR is most sensitive to variations in metal prices and mined grades and least

sensitive to operating costs. Simulated variations in the potential expected recoveries of

payable metals show limited sensitivity but the project will be economically vulnerable to a

variations in metal recovery rate.

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Figure 1.1 Pre-tax IRR Sensitivities

Figure 1.2 Project NPV Sensitivities

The following figures isolate the sensitivity of the project to variations in the price of gold only.

0

5

10

15

20

25

30

35

40

45

50

-15% -10% -5% 0% 5% 10% 15%

IRR

(%

)

Percentage Change In Variable (%)

Metal Prices Capital Costs Operating Costs

Mined Grade Metals Recovery

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Figure 1.3 Project IRR Sensitivity to Gold Price

Figure 1.4 Project NPV Sensitivity to Gold Price

0

5

10

15

20

25

30

35

40

45

50

-15% -10% -5% 0% 5% 10% 15%

IRR

(%

)

Percentage Change in Metal Prices (%)

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1.1 Conclusions

Based on the study results, conclusions are:

1. The project provides positive and robust returns.

2. Longstreet is a small deposit which can be developed for production at reasonable

cost in a near-term horizon, provided regulatory approval and permits are acquired.

3. The mined grade of ore is an important variable for the success of the operation as is

mining cost. Operating management efforts during mine production must be

focussed on these two parameters.

4. The Project is most sensitive to variations in the price of gold and variations in the

mined grade of mineralized material.

5. The economics of the project would be improved with the discovery and exploitation

of economically viable satellite deposits. Once the capital investment has been

repaid by the Main Zone the operating profits from other deposits would enhance

the Project cash flow.

6. Water sourcing is the largest technical risk factor, particularly to capital expenditures

and operating cost estimates. Ideally a well source will be identified and thus avoid

the added cost of piping water to the site from Five Mile Spring on Clifford Ranch,

currently the nearest identified water source, 12 miles from the Project site. It is not

known if Five Mile Spring produces adequate volume nor if the owner of Clifford

Ranch would agree the sale of water from the spring. Preliminary investigations have

identified two potential sources of well water; one mile to the NE in the Monitor

Range and approximately five miles east in Stone Cabin Valley.

7. AMPL has reviewed the permitting requirements of the US Forest Service, The

Bureau of Land Management and the Nevada Division of Environmental Protection

and estimates that, without objection during the public disclosure period of

permitting, the Longstreet Project will require from two to four years to secure the

permits required to begin constructing and operating the mine.

1.2 Recommendations

Based on the results of this Scoping Study, recommendations follow.

Geology

Recommendations for the next phase of mineral resource estimation include:

1. Consider a drilling program to explore mineralization at depth and test the nearby

Central Ridge mineral occurrence.

2. Consider further drilling to better understand the transition zone between oxide &

sulfide to determine the maximum extent of leachable gold mineralized material.

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Mining

1. Geotechnical work for open pit slope angles optimization is recommended.

2. Firm quotations from qualified local mining contractors is advised for the next phase

of study.

Heap Leaching & Processing Plant

1. Conduct column test work on the oxide adit material to test the mineralogical variability of the deposit.

2. Conduct column leach tests using finer material in conjunction with high pressure rolls i.e. P80 ¼-inch (6.3mm) in order to maximize silver recovery.

3. Reduce column leach time to 60 days of leaching based on gold and silver recovery testwork results.

4. Conduct column tests using site water as opposed to laboratory tap water in order to determine the effects of site water on leach kinetics.

5. A HPGR (high pressure grinding rolls) evaluation should be considered to investigate improved silver recovery on the master blend composite ore (generic tests show that HPGR use often leads to the formation of micro cracks in the ore which may improve silver leaching kinetics). This would require a Static Pressure Test (SPT) to be performed.

6. Load/permeability tests are recommended on column leach residue samples to confirm permeability under compressive loading.

Infrastructure

1. A hydrological study is recommended to identify proximal water sources of adequate

volume to sustain the Longstreet operation.

Environment and Permitting

1. Initiate baseline studies as soon as possible as a precursor for applications for

permits to construct and operate the Project.

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2 INTRODUCTION

Star Gold Inc. (Star Gold) engaged A-Z Mining Professionals Limited (AMPL) of Thunder Bay, Ontario, Canada, to undertake a scoping study of the Longstreet Gold Project, located 50 miles NE of the town of Tonopah in Nye County, Nevada, USA. This study uses the mineral resources estimate prepared by Agnerian Consulting Limited in December 2013 and reported in the NI 43-101 report entitled “Technical Report on the Longstreet Gold-Silver Property, Nevada”, dated December 15, 2013. For this study, the Qualified Persons to have visited the site include Mr. Joe Kantor (Geology),

Mr. Reinis Sipols (Mining Engineer and Environmental Specialist) and Mr. Dan Peldiak

(Metallurgical Engineer), the latter on a prior study when employed by Coffey Mining.

Star Gold has made every effort to provide access to the property and ensure that all current

and historical technical data, was available for AMPL to review. In addition, Mr. Richard Kern, a

Reno Nevada, USA, based geologist with many years association with Longstreet and vendor of

the property, was most helpful in providing access to historical documents, maps and assays as

well as hosting a site visit. AMPL appreciates the assistance of all those involved.

All measurement units in this report are Imperial unless otherwise stated.

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3 LOCATION & HISTORY

The Longstreet Project is located in a historically prolific region of mineral production in Nye

County, Nevada, known as Walker Lane. Walker Lane hosts the well-known deposits of Round

Mountain, Mineral Ridge, Bell Mountain and Bullfrog, all current or previous producers.

The project is located approximately 170 miles northwest of Las Vegas and approximately 50

miles northeast of Tonopah, a town of approximately 2,500 people and the seat of Nye County,

in west-central Nevada. The northeast-southwest oriented property is situated within the

McCann Canyon and Georges Canyon Rim, covering 7-1/2 topographic quadrangles, and

extends approximately two miles along strike within the Monitor Range. The approximate

geographic coordinates of the central part of the property are 38° 22’ 00”N Latitude and

116° 40’ 00”W Longitude (Figure 3.1). Figure 3.2 shows an aerial view of the property from

satellite imagery.

The deposit has been known for many years and the property explored on numerous occasions.

Exploration work on the property has included pits, core drilling, RC drilling, an inclined shaft,

three adits and limited underground vertical raising. The deposit has never been mined on a

commercial scale. Figure 3.3 shows the surface plan of the property and Figure 3.4 shows the

underground accesses locations in section view.

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Figure 3.1 Longstreet Gold Project Location

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Figure 3.2 Longstreet Gold Project Satellite View (north to top of page)

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Figure 3.3 Longstreet Gold Project - Main Zone

(red line = Resource estimate pit shell, after Agnerian)

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Figure 3.4 Cross-section through Upper Adit

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4 GEOLOGY

4.1 Regional Geology

The Longstreet property is located in the Monitor Range, a westward-tilted fault block elevated

by normal faults and is part of the Basin and Range Province. The ranges are topographic highs

surrounded by thick alluvium-filled valleys. The southern Monitor Range consists mainly of

Tertiary age volcanic rock related to the Big Ten Peak volcano and a nearby unnamed 29 Ma

caldera intruding and overlying Paleozoic sedimentary and metamorphic rocks.

4.2 Property Geology

The Longstreet deposit is a fossil hot-spring related gold-silver deposit hosted in Oligocene-age

weakly to moderately welded rhyolite volcanic rock (Unit Tat). Mineralization is controlled by

the east-west striking, 40° to 55° north dipping Adit Fault vein and northwest and west-

northwest striking steeply north dipping sheeted veins and fractures. The bulk of the gold-silver

mineralization, termed the Main Zone, is in steeply dipping multiple sheeted vein sets parallel

or sub-parallel to the moderately dipping Adit Fault. Mineralization is associated with quartz-

adularia, clay and pyrite with oxide, transition and sulfide zones. The end of the mineralization

cycle is expressed as a volcanoclastic and sinter member (Unit Ts) with local anomalous gold-

silver values. A barren, post-mineral volcanic unit (Trt) overlies the volcanoclastic and sinter

member. Intruding into the Tat unit are unaltered rhyolite intrusive dikes (Trp) which may be

the feeder zones to the overlying, younger post-mineral volcanic rocks.

Peripheral to the Main Zone deposit are eight target areas with anomalous structure, alteration

and/or gold-silver geochemistry. Nearly all have been explored to some extent with either

angle and vertical air track drill holes or Reverse Circulation (RC) drill holes.

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5 RESOURCES

The basis of this Scoping Study is a resource estimate (Table 5.1) prepared by Agnerian

Consulting Limited and reported in National Instrument (NI) 43-101 format (a Canadian

reporting standard for the mineral industry). The resource estimate of record is contained in the

report, Technical Report on the Longstreet Gold-Silver Property, Nevada, dated December 15,

2013.

Table 5.1 Mineral Resources - Longstreet Gold Project

Mineral Resources – Star Gold Longstreet Project, Nevada

Category Tonnes Grade (g/t Au)

Contained Ounces Au

Grade (g/t Ag)

Contained Ounces Ag

Indicated 4,394,000 0.64 90,900 15.64 2,210,000 Inferred 304,800 0.48 4,750 14.56 142,700

From: Agnerian, Dec 2013, p65, Table 14-1

AMPL notes that the Agnerian resource estimate applied economic constraints and a

preliminary pit shell, reporting only those resources which were contained by his pit shell.

AMPL does not consider this to be best practice in that economic constraints, which are a key

variable, may inaccurately portray the totality of the mineral inventory at the resource

estimation stage.

AMPL was provided with the Agnerian geological block model, which includes the total mineral

inventory at Longstreet, and its mining engineer applied current economic parameters to

develop an optimized pit shell and thereby estimate a mineable resource.

5.1 Resource Comparison

Star Gold had commissioned two separate 43¬101 reports from different firms for the

Longstreet Property Main Zone. Each study developed different block models based on

different methodologies. The first is that of Agnerian dated Dec 15, 2013 which we have

referenced and which forms the basis of this scoping study. The second report is authored by

Paul Noland titled “Longstreet Project, Nye County, Nevada Revised Technical Review and

Resource Estimate February 16, 2014 (hereafter called the “Noland Report”). Star Gold

requested an analysis of both reports to enable this scoping study to choose the more

technically proficient and appropriate block model.

5.2 Block Model Review

Star Gold requested a Qualified Persons (QP) geologist from AMPL conduct an independent

review of geology block models created independently of each other by Agnerian Consulting

Limited and Noland Engineering, for the Longstreet deposit. The following is an evaluation of

each model, performed by Alan Aubut, P.Geo.

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5.2.1 Database

The quality of any resource estimate is at its most fundamental level dependent on the quality

of the data used. AMPL has compared the two block models and the methods used to estimate

mineral resources therein. A summary of the key comparisons is provided in Table 5.2.

Table 5.2 Comparison of Agnerian and Noland Block Models

Elements of a Good Resource Estimate

Item Agnerian Noland

1 Mineral Envelop captures just those samples that are part of the mineral

system.

Yes No

2 Block Size is representative of envisioned Smallest Mining Unit Yes No

3 Complete Variography so that samples can be used to measure the range

(distance) over which they show some correlation. These distances can

then be used to select related samples when doing the actual estimation.

Yes No

4 Select an estimation method that is not prone to conditional bias (grade

too high and tons too low), typically a variant of Kriging.

Yes No

5 Also do a parallel model, preferably Nearest Neighbour, to establish a base

line for the global mean which can then be used as a first pass during

validation as all methods should produce essentially the same global

mean.

Yes No

6 Report Resource using only a cut-off as to do otherwise is in effect defining

Reserves which then become invalid if any of the economic criteria have

changed or are incorrect.

No ?

The source database was not provided to AMPL, therefore no specific comments can be made,

but two things should be reviewed that can have unsuitable impacts on a resource estimate:

1) Whether to use absent data rather than setting all unsampled intervals to 0 (zero)

grade; and

2) Use of grade capping.

For 1) where unsampled intervals are present in the estimated mineral domain, the block

estimate is biased high, as all samples with absent data will be ignored. The impact is that block

grades involving such samples would be too high, thus possibly over estimating the grade.

While grade capping is commonly used in the gold mining industry to try and compensate for

the nuggety nature of that type of mineralisation, in actual practice it has no statistical validity.

Thus it should not be used as it can artificially depress the true grades. One Geostatistical

method used to overcome this problem and best suited for highly skewed distributions (a low

grade peak with a long high grade tail) is Multiple Indicator Kriging. Note that this is the method

that was used by Agnerian and is considered Industry Best Practice.

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5.3 Mineral Wire Framing

A critical element of any resource estimate is the definition of the mineral domain through the

use of wire frames. These wireframes need to isolate samples that can be considered part of

the mineralised system from samples that are not. To do otherwise introduces a serious bias.

In the case of the Agnerian model, there are five domains, one for the Main Zone and four for

the Footwall Zone. All domains are reasonably well constrained and eliminate large areas

barren of mineralisation.

On the other hand, the Noland model used a single domain that is far too large and

encompasses everything from no grade to high grade. This introduces several serious biases

including:

Allowing the spreading of grade far beyond where gold would reasonably be expected to

occur.

Allowing areas to be under estimated because too much low grade material has been

used in any one particular block estimate.

Sample sections showing the two models are presented in Figures 5.1 and 5.2.

Figure 5.1 Sample section of Agnerian geological model

Note the domains constrain the model to just those portions that are mineralised.

______________________________________________________________________________________________


Figure 5.2 Sample section through Noland model (same section)

Note the single domain which does a poor job of constraining the model and allowing trends different than those demonstrated by the Agnerian model as shown by the heavy gray lines. Also note how, in general, the mineralized areas of the Noland model are larger.

While the Agnerian domains can be considered up to Industry Standards, the Noland single

domain cannot.

5.4 Sample Capture and Compositing

The use of 5 foot drill core composites for sampling and assaying is quite reasonable. The

Agnerian domains captured only those areas with mineralisation, except where more recent

drilling has taken place. Due to its large size the Nolan Domain rejected only a minor number of

samples.

5.5 Variography and Variogram Modelling

Variograms are the measure of similarity between samples with increasing distance from the

drill intercept point. As such, variograms allow the determination of what samples are to be

selected based on the samples themselves, rather than an arbitrary number. Likewise,

variogram models provide a means of calculating sample weights as determined by the samples

and their innate relationship between one another. The use of variograms therefore provides

the least biased method of determining sample weights when doing resource estimation.

Kriging is the one estimation method that takes full advantage of these characteristics of

variograms and as such is considered Industry Best Practice.

______________________________________________________________________________________________


The Agnerian report provided examples of variograms for both gold and silver. Several

observations can be made:

- The variograms are well formed indicating that the drilling density and sampling was

adequate for generating reliable variograms, thus confirming there is enough data to

allow the use of Kriging.

- The variogram models are good with a reasonable nugget effect and all variograms are

indicative that Industry Standard Practice, at least, has being followed.

The Noland modelling did not use variography, instead relying on “known geologic trends at the

project site, mirroring parameters of a decade of hand-drawn resource calculations”. While this

cannot be considered a “Fatal Flaw” (an error that in itself would invalidate the estimate), it

does not meet the threshold of current Industry Standard Practice.

5.6 Block Model Definition

A term commonly applied to each individual block in a block model is Smallest Mining Unit, or

SMU. The block modelling practitioner, with advice from a mining engineer if available, must

make a decision as to what type of mining will likely take place, keeping in mind the amount of

selectivity anticipated.

The Agnerian model uses blocks that are 20 feet cubed and in line with what would be

considered the smallest mining unit for open pit mining.

While not considered a fatal flaw, the SMU used by Noland is considered too small as 10 foot

square blocks are more appropriate for underground mining where these dimensions would

approximate a single blast round. As a result the estimated grade by Noland, especially for the

higher grade blocks, is likely to be significantly higher than will be experienced in practice as the

deposit is likely to be mined by open pit methods.

5.7 Estimation Search Criteria

Any block modelling method is only as good as the search criteria used in the estimation

process. A potential source of bias is the use of samples from too few drill holes. This can be

alleviated by using an octant search methodology (the search ellipsoid is divided into 8 parts, or

“octants” based on the 3 primary planes) and then making sure samples from enough octants

are used to ensure suitable distribution in 3D space.

The Agnerian and Noland searches used similar ranges with the former being determined from

variography. Both used a nested search approach with the latter involving doubling of the

initial search ranges.

In the case of the Agnerian model too few samples were used. Only samples from a maximum

of two holes were used and a maximum of eight samples in total. The octant search method

was not used thus allowing results to be biased to isolated parts of the block.

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At the time of writing of this review no information was available for the actual search strategy

used by Noland, other than the aforementioned search ranges.

5.8 Statistical and Visual Checks of Model

Industry Standard Practice is to conduct two estimates using two different estimation methods

as the global mean should be essentially the same no matter what method is used. A good

starting point for example is to create a Nearest Neighbour model. This method relies on using

the nearest sample for estimation and is very similar to the old polygonal method. While locally

biased it does provide a reliable source for determining the global de-clusterised mean, yet is a

method where it is difficult to make a mistake.

The Agnerian model was estimated using Kriging and Nearest Neighbour and the mean grade

for each is within 5% of one another, as would be expected, providing confidence in the

estimate.

Noland, on the other hand, only created an Inverse Distance model so there is nothing to

compare the results of this model with to ensure it meets the fundamental requirement of

maintaining the global mean. Further, it would be misleading for the reviewer to compare the

mean of the model with that of the drill holes as the latter have significant built in bias due to

clustering (too many samples in isolated pockets, typically in high grade areas).

A standard check is to make sure the block grades are reasonably spread out and honour the

grade values ascribed to the relevant drill holes. The Agnerian block model matches the

informing holes reasonably well. The Noland model shows significant bias in areas due to the

unconstrained nature of the domaining, as well as a result of using the Inverse Distance method.

These two issues combine to form “bulls eyes” in areas with poor sample availability, with the

high grade core much larger than can be supported by the available drill data (see Figure 5.3).

______________________________________________________________________________________________


Figure 5.3 Detailed section through Noland geologic model

The section in Figure 5.3 shows: 1. poor correlation with the informing drillhole; 2. does not honour the established geologic trend due to the use of too large of a

domain; and 3. shows a serious bias as illustrated by the “bulls eye” effect common with Inverse

Distance and poorly sampled areas.

5.9 Resource Reporting

A resource estimate should be the best estimate possible using the available data. Various

mineral resource reporting codes around the world specify that a mineral resource must “have

reasonable prospects for economic extraction”. The key words here are “reasonable” and

“prospects”:

Reasonable – as much as is appropriate or fair; moderate.

Prospect - the possibility or likelihood of some future event occurring.

Note the vagueness of the two definitions! And there is good reason as there are many factors

that typically are not known at the time that an estimate is generated that will determine

whether or not economic extraction is possible. These factors include metallurgy, access to

infrastructure and metal prices, all of which can and will change with time. But a resource

estimate should be a snap shot that can stand on its own merits.

______________________________________________________________________________________________


It is said that mines are made and not found. The same cannot be said of the resource as a

mineral resource can only be found, not manufactured; although the confidence in the resource

may be improved through the application of direct testing methods such as diamond drilling

and bulk sampling.

Thus for a mineral resource to be useful in the future, factors that are beyond the control of the

person who generates the estimate should not be applied to defining the resource – factors

such as metal price and metal recovery. To do otherwise introduces a serious bias in that the

quantities reported will be incorrect if any of these factors change.

Agnerian has used an approach that introduces a serious time sensitive bias (though also used

by other consultants in open pit resource estimates). The Agnerian report is based on the

application of a pit “shell” design that in itself is based on metal prices and recoveries that will

most certainly change. The resulting “in-pit” resource thus gives an unclear measure of the

entirety of the mineralization – information that is very important if there is a change in

commodity prices or gold recovery techniques.

This review did not encounter information in the Noland resource report that would indicate

the methods employed, or constraints applied, for the resource estimate methodology,

therefore no comment is made thereon.

5.10 Conclusion

The Agnerian model in general applied Standard to Best Industry Practice methodology. While it

is deficient in some areas, specifically the search strategy employed, none of the deficiencies

are considered Fatal Flaws that would invalidate the estimate.

A number of serious issues were found in the methodology used and the application of block

modelling techniques by Noland. These issues include lack of the application of variography, use

of Inverse Distance estimation and improper domaining. While none of the issues can be

considered a Fatal Flaw in that it resulted in a gross error of the estimate, they do result in a low

confidence in the estimate which would relegate the mineral resources to the Inferred category,

in the opinion of AMPL.

5.11 Geology and resource estimate recommendations It is recommended that any mine design work be conducted using only the Agnerian model:

1) The table of resources should not use any control other than a grade cut-off, as to do

otherwise introduces a time sensitive bias.

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6 GEOTECHNICAL AND HYDROGEOLOGY

6.1 Regional Hydrogeology

The Longstreet Mine is located in the Monitor Range of the Great Basin section of the Basin and

Range physiographic region. The Basin and Range is largely an arid region encompassing the

majority of the western United States; its topography is characterized by alternating narrow

faulted mountain ranges and flat fault bound valleys (basins). The Great Basin, as a section of

the Basin and Range region, follows the same topographic characteristics with notable internal

surface and subsurface hydrologic drainage.

The aquifer system within the Great Basin generally comprises aquifers in unconsolidated

alluvial fill, sedimentary and volcanic deposits in fault bounded basins, and in various bedrock

lithologies of the mountain ranges that drain into the separate basins. The mountain range

bedrock units often underlie the basins. The basic hydrogeologic model is illustrated in Figure

6.1. The mountain range consists of consolidate bedrock with limited unconsolidated alluvial fill.

The bedrock in the mountain ranges generally is less porous and permeable rocks compared to

the basins’ bedrock. These rocks are characterized by fractured flow conditions. The resulting

lower permeability impedes groundwater flow and the fractured flow conditions in many cases

limit the groundwater volume available as a resource. (3)

One of the limiting factors for water availability in the Great Basin is the low water recharge to

the local aquifers due to the limited precipitation in the area (3,4). The Great Basin resides in

the rain shadow of the Sierra Nevada Mountains thus precipitation is limited and irregular; the

least precipitation occurs in the valleys and the greatest in the mountains (2, 3, 5). Winter

precipitation generally consists of snow, and summer precipitation is characterized by localized

high intensity rain (4). Geologic evidence and recorded history indicate the intense rain storms

may result in flooding of the major rivers and the “dry” washes (4), and due to the arid

conditions the evaporation rate is high. Precipitation that does not evaporate either percolates

into the subsurface or moves as surface runoff into the valley basins (2, 5) thereby the basin

aquifers are recharged. As illustrated in Figure 6.1, both the surface runoff, and the

groundwater in the ranges flow into the valley basins, therefore the valleys are the best sources

of water.

6.2 Local Hydrogeology

Limited information is available regarding the actual water resources that exist within the

Monitor Range where the Longstreet project is located. Three springs are mapped on the

eastern edge of the Monitor Range: the Painted Rock Spring, Side Hill Spring, and Four Mile

Spring. All of the springs are relatively close to the Longstreet deposit. They are located at or

near the topographic transition between the Monitor Range and the Stone Cabin Valley. The

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Figure 6.1 Great Basin Hydrogeology

existence of springs indicate potential exploitable groundwater in the bedrock, and/or the

alluvium of the ephemeral streams that flow east out of the Monitor Range. The volume of

water flowing from the springs, and any seasonal variation, is not known. The Side Hill Spring is

the closest; it is located approximately 1.5 miles east of the Longstreet mine.

The greatest potential source of groundwater is Stone Cabin Valley. The valley has a drainage

area of 961 square miles with a net recharge to the basin of 16,000 acre feet of water per year.

In 1962, the cumulative ‘loss/use’ via evapotranspiration and reclamation was estimated to be

2,000 acre feet per year (2). Since 1962 no significant development has occurred to alter this

‘loss/use’ estimate. This difference between recharge and discharge rates indicate this

groundwater resource could supply a substantial amount of water without significantly

lowering the groundwater levels or negatively affecting existing groundwater use within the

valley.

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Stone Cabin Valley is principally utilized as livestock range. The Clifford Ranch is the only

identified active ranch in Stone Cabin Valley. The ranch is located approximately 12 miles south-

south east of the mine site. Five Mile Spring is located at this ranch.

At Mud Lake, Stone Cabin Valley surficially drains into Ralston Valley. A well field exists in the

Ralston Valley supplying water to the City of Tonopah. The water from this field is transported

the 15 miles to Tonopah by pipe.

6.3 Mine Site Hydrogeology

No information is currently available regarding the presence of water at the mineral project site.

Preliminary interpretation regarding the potential availability of water can be predicated with

the existing geological maps created by previous mine owners. The maps indicate that the

Longstreet property is underlain predominantly by Oligocene Epoch moderately–to poorly-

welded tuffs with common lithic and pumice fragments. Four lithologic units have been

described at the Site (1):

“Welded Ash Flow Tuff (Tat) -This rock is buff to grey, and contains <10% fine-to medium-

grained quartz phenocrysts, 15% fine-to medium-grained feldspar phenocrysts, 5% to 15%

medium to coarse-grained pumice, and 5% to 20% other “exotic” fragments in an aphanitic

groundmass. The rock displays horizontal bedding and may be up to 3,000 feet thick. It exhibits

pervasive hydrothermal alteration consisting of argillic alteration (bleaching and clay mineral

development), silicification (quartz flooding and/or networks of numerous quartz veinlets), and

potassic alteration (adularia in quartz veinlets). Supergene limonitic and goethite alteration

overprint the hydrothermal alteration.

Rhyolitic Porphyry Dike (Trp) - Rhyolitic porphyry dikes of various orientations intrude the Tat

unit, and may be associated with the heat source of the mineralizing fluids at Longstreet.

Siliceous Sedimentary Rock (Ts) - A thin unit of white, yellowish and grey, volcaniclastic and

siliceous rock (including sinter) intermittently overlies the Tat unit. Silicic alteration is evidenced

by sheeted quartz veins.

Welded Tuff (Trt) - Black to brown, strongly welded tuff occurs along ridges and overlies the Tat

and Ts units. This unit is 330 feet to 400 feet thick and has a distinctive thin (approximately 3

ten-feet) vitrophyre zone near its base.”

This indicates the welded tuffs have a limited capacity to store water (porosity), or allow water

to flow (permeability). The welded strength of the tuff affects porosity and permeability: the

greater the welding, the lower the porosity and permeability. Therefore the strongly Welded

Tuffs, by definition, have low porosity and permeability (6). The Welded Ash flow tuff and the

Welded Tuff are interpreted to be dense and relatively impervious rocks. The Siliceous

Sedimentary Rock offers a potential porous media for groundwater however its thickness may

limit the volume of water that it can store. No information is available regarding the presence of

water in these lithologic units.

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Suggested locations for potential well sites are identified in Figure 6.2.

Figure 6.2 Water well search location recommendations

Well Location Option 1 is approximately 0.7 miles from the Longstreet Project. Water is

expected but volume and sustainability is unknown. The location is chosen because the surface

drainage is at the intersection of three sub-watersheds.

Well Location Option 2 is approximately 5 miles from the Longstreet Project. Water is expected

with the necessary volume and sustainability. This location is at the valley centre and assumed

to be the best location for water production. Other locations can be identified between this

site and the entrance to Windy Canyon that could potentially supply the volume and long term

sustainability requirements of the Project.

6.4 Hydrogeology conclusion and recommendation

Water sourcing is the largest technical risk factor, particularly to capital expenditures and

operating cost estimates. Ideally a well source will be identified and thus avoid the added cost

of piping water to the site from Five Mile Spring on Clifford Ranch, currently the nearest

identified water source, 12 miles from the Project site. It is not known if Five Mile Spring

produces adequate volume nor if the owner of Clifford Ranch would agree the sale of water

______________________________________________________________________________________________


from the spring. Preliminary investigations have identified two potential sources of well water;

one mile to the NE in the Monitor Range and approximately five miles east in Stone Cabin Valley.

A hydrological study is recommended to identify proximal water sources of adequate volume to

sustain the Longstreet operation.

6.5 Geotechnical

Geotechnical descriptions of the rock at the Longstreet Property were not available for review.

______________________________________________________________________________________________


7 MINING

7.1 Potentially Mineable Mineral Resource – Pit Optimization

The potentially economic mineral resources were defined as those blocks falling within an

optimized pit shell derived from the economic parameters shown in Table 7.1. The unit costs

used in the pit optimization process were based on preliminary estimates and general

knowledge of mining, processing and general and administration costs for similar type

operations. The pit optimization was conducted using Mintec MineSight Economic Planner

2.60-00 pit optimization software.

Table 7.1 Floating cone pit optimization parameters

Parameter Value used in Floating Cone Pit Optimization

Gold price $1,350/ troy ounce

Gold recovery 86%

Gold transport and refining charge $4 /troy ounce

Silver price $24

Silver recovery 15%

Waste mining cost $3.00 / ton

Mineralized material mining cost $3.50 / ton

Heap leach crush and place cost $3.50 / ton

Processing cost $3.65 / ton

General and Administration cost $2.01 / ton

Assumed pit slope angle 50°

Base cone radius 40 feet

The potentially mineable mineralization was determined using a breakeven cut-off where

revenue is equivalent to marginal costs. The $9.16/ton breakeven cut-off, derived from the

sum of the estimated processing and G&A costs, does not include mining costs as all material

contained within a shell is considered mined and sent either to the waste dump or the leach

pad.

The 50° pit slope angle has been assumed and is based on the experiences of other mining

operations in the region. There may be an opportunity to steepen the pit slope but this would

need to be demonstrated by a geotechnical investigation and assessment as part of future

studies.

The geological block model developed by Agnerian Consulting Ltd. as reviewed by A. Aubut of

AMPL was found suitable for use in this internal scoping study – see also Mining Conclusions

and Recommendations.

The in-pit potentially mineable mineral resources estimate is shown in Table 7.2 and may be

materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing

______________________________________________________________________________________________


or other relevant issues. The mineral resources estimate takes geologic, mining, processing and

economic constraints into account, are confined within a pit shell, and are classified in

accordance with CIM Definition Standards for Mineral Resources and Mineral Reserves.

Table 7.2 In-pit Mineral Resource Estimate

Mineral Resource Category

Short Tons Au (oz / ton)

Ag (oz / ton)

Indicated 3,982,339 0.0219 0.5207

Inferred 225,455 0.0189 0.6076 1. Mineral Resources which are not Mineral Reserves do not have demonstrated economic viability.

2. The quantity and grade of reported Inferred Resources in this estimation is uncertain in nature and there has

been insufficient exploration to define these Inferred Resources as an Indicated or Measured mineral resource,

and it is uncertain if further exploration will result in upgrading them to an indicated or measured mineral

resource category.

3. The mineral resources are reported within the optimized pit shell that was used to assess reasonable

prospects of economic extraction. The mineral resources estimate excludes external dilution and mining

losses.

Plans and sections of the pit shell are shown in Figures 7.1 to 7.5.

The in-pit potentially mineable mineral resources estimate was prepared using Mintec

MineSight Economic Planner pit optimization software, and the geological block model for the

Longstreet Star Gold deposit received on January 6, 2014.

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Figure 7.1 Pit shell aerial view

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Figure 7.2 Pit shell longitudinal section, 13,935,430E

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Figure 7.3 Pit shell longitudinal section 13,935,430E with topography

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Figure 7.4 Pit shell cross-section 13,934,000N

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Figure 7.5 Pit shell cross-section 13,935,145N

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7.2 Mining Method

The topography at Longstreet is advantageous for open pit mining in that there is very little

waste rock that must be stripped prior to the commencement of production mining operations.

Year 1 of the mining schedule will deliver the scheduled ROM tonnes and excavate necessary

waste rock. There is no requirement for a capital allocation to pre-strip the planned open pit

mine.

Pre-production work would include establishing the main haul road to the heap leach and

surface road to mine facilities including explosives magazines.

The open pit would be mined using conventional mining equipment and technologies.

Mineralized material and waste rock would be blasted, excavated, loaded and hauled to either

the waste rock management area or the heap leach crusher. It is assumed that a contractor

would develop and operate the pit, crush the mineralized material, place and spread the

mineralized material on the leach pad and prepare the surface of the stacked material using a

ripper. The type of equipment used would depend upon the contractor’s equipment

preferences and available fleet. It is envisaged that 20 foot benches would be used in the pit

and that the contractor would use conventional mining equipment such as a track-mounted drill,

hydraulic excavator, wheel loader, 40-ton class trucks and bulldozers. It is expected that the pit

would operate 350 days per year and the mining fleet sized accordingly.

It is assumed that the pit would be dry and that a conventional diesel-powered pump would

only be required from time to time to dewater the pit sump.

This study considers that the mining contractor would supply its own equipment and shop and

that the pit access road and minor pre-stripping would be done concurrent with the

construction of the leach pad.

It has been assumed that the mine owner would manage the project and provide technical

services.

7.3 Mining Schedule

The mine schedule is based on the optimized pit plus mining dilution (5%) and losses allowances.

The total tons of material that would be mined from a designed pit would be expected to add

marginally to the strip ratio. The impact on the economics of the operation are minimal.

Mining activities have been planned and scheduled to address pre-stripping of waste rock, ore

and waste rock mining throughout the life-of-mine (LOM). The design team selected a run-of-

mine (ROM) ore production rate of one million short tons per year, forming the basis for the pit

design and the processing systems. The mineral resources incorporated into the pit shell are

adequate for 4.4 years of ROM production. The LOM strip ratio is a favorable 0.7 tons waste : 1

ton ore.

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The mine schedule shown in Table 7.3 makes use of Inferred Mineral Resources. The scoping

study is preliminary in nature and includes Inferred mineral resources that are considered too

speculative geologically to have the economic considerations applied to them that would

enable them to be categorized as mineral reserves, and there is no certainty that the results

indicated in the scoping study will be realized.

Table 7.3 Mine Schedule

Item Year LOM

-1 1 2 3 4 5

Leach pad feed

(k ton)A

Au (oz / ton)B

Ag (oz / ton) B

1,000

0.0217 0.5254

1,000

0.0217 0.5254

1,000

0.0217 0.5254

1,000

0.0217 0.5254

418

0.0217 0.5254

4,418

0.0217 0.5254

Waste rock

(k ton)

697

697

697

697

291

3,081

Strip ratio 0.7

A Leach pad tonnage includes a net 5% allowance for mining dilution and mining losses. B Average grade assumed delivered to leach pad over LOM.

7.4 Mine Closure

The regulatory requirements for mine closure and site reclamation are well established in

Nevada. A tentative permanent closure plan would need to be submitted at the time of the

application for a Water Pollution Control Permit, and the final permanent closure plan would

need to be submitted two years before the anticipated closure of the site. The final closure

report must be submitted to the Nevada Division of Environmental Protection, Bureau of

Mining Regulations and Reclamation following the completion of closure to demonstrate that

the Waters of the State will not be degraded, and propose the post-closure monitoring program

to regulators.

The Longstreet Project is still at the conceptual stage and a tentative permanent closure plan

has not yet been developed. The plan would be expected to encompass but not be limited to

the collection and responsible treatment and/or the permitted disposal of process solutions,

reagents and hazardous wastes, used oil, and non-hazardous materials and wastes; the orderly

removal and/or demolition of process equipment and buildings; closure works to ensure that

the pit and stockpiled mine materials are left in physically and chemically stable conditions; the

access road would be reclaimed; controls would be put in place to prevent inadvertent access

into the mined-out pit; run-on interception and diversion ditches; contact water interception

and management; dust control measures; and other measures to protect human health and the

ecology over the long term; and a monitoring program to provide data to demonstrate the

______________________________________________________________________________________________


effectiveness of the closure works and site reclamation. The cashflow model for the project

includes a closure and reclamation cost allowance.

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8 METALLURGICAL TESTWORK and MINERAL PROCESSING

8.1 Historical metallurgical sampling and test work

A pre-feasibility study was conducted on the Longstreet project in 1988 by Mining Engineering

Services. Metallurgical test work was performed in support of the pre-feasibility study, which

consisted of bench scale bottle roll tests on 10 samples composited from 31 drill hole

composites. These samples were crushed to -10 Mesh (Tyler Series) and used in the bottle roll

tests. Results from the test work indicated that a gold recovery between 82.1% to 87.2% could

be achieved along with a silver recovery which ranged between 28.3% to 57.9%.

In addition, a large bulk sample was collected from three surface pit sites and four underground

sites. It is not known as to how the underground samples were collected. The bulk sample was

screened and split into six individual samples for further metallurgical testing. Test work was

carried out on +3 inch (+76mm) material for bucket tests, -3 inch material for column tests and -

1/4 inch (6.35mm) material for column tests. Test results indicated that gold recovery for the +3

inch material ranged from 50% to 63%, gold recovery for the -3 inch material ranged from 68%

to 87%, and gold recovery for the -1/4 inch material ranged from 86% to 90%. Results are listed

in the Table 8.1.

Table 8.1 Metallurgical Testwork Results, (c. 1988)

For material similar to that tested, Kappes, Cassiday & Associates (KCA) estimated field heap

leach recoveries to be 85% for gold and 20% for silver using ¼-inch material.

In April of 2012, Paul D. Nol and KCA published a Technical Review and Resource Estimate for

Star Gold, in which they reported results obtained from a previous test work program by Harron

(2003) and MDA (1988). The test program involved compositing numerous oxide drill intercept

cuttings in which bottle roll tests were performed on 10 samples. Average gold recovery results

for -10 Mesh samples were 85.4% gold and 37.9% silver recovery in 72 hours. KCA then

conducted column tests on three samples to test the responses of low, medium and high grade

material from underground. After crushing to -3/4 inch (19 mm) the samples averaged 82% gold

Size Days Leached

Au g/t Ag g/t Au Ag

+76 mm(s) 44 0.342 11.51 63.6 <1.0

+76 mm(u) 44 0.995 41.06 50.0 4.6

-76 mm(u) 45 1.275 37.01 87.8 10.9

-76 mm(s) 45 0.778 16.48 68.0 15.1

-6.35 mm(s) 42 0.684 14.93 86.4 25.0

-6.35 mm(u) 42 1.026 33.90 90.9 23.9

(s) -surface

(u)- underground

% RecoveryCalculated Head

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and 29% silver recovery. Crushing to -6 mesh 0.132 inches (3.6 mm) increased recovery to 93%

for gold and 52% for silver. According to the test work conducted those are the expected

recoveries for an open pit heap-leach operation at Longstreet. The data was generated 25 years

ago, on underground samples only.

KCA also conducted agitated cyanide tests on pulverized material and obtained 92% gold and

81% silver recovery. These are the recoveries expected in a conventional mill utilizing a fine

grind.

Column leach tests were also conducted on behalf of Bacon-Donaldson Engineering on -2 inch

(50 mm) material. Recoveries varied from 85% to 90% for gold and 9% to 28% for silver, with

underground samples being more amenable to leaching than surface samples. It appears the

oxide zone of the Main deposit has reasonable leaching characteristics for gold although silver

recovery is poor.

The current resource estimate identifies two main types of mineralization near surface: “oxide”

mineralization associated with limonitic pseudomorphs of pyrite, and “sulphide” mineralization

associated with disseminated pyrite. The oxide mineralization was the focus of the 1988 pre-

feasibility study which achieved recoveries of approximately 90% for gold and 25% for silver

based on column leach tests utilizing a feed size of ¼ inch (6 mm). Later tests on coarser oxide

material of -2 inches (50 mm) resulted in recoveries of 90% for gold and 28% for silver indicating

high gold recoveries can be achieved using coarse material. These results contradict current

findings as results obtained in lab bottle roll tests indicate that crush size does not have a

significant impact on silver recoveries and that current adit samples used in column tests

exhibited inferior leaching kinetics as compared to surface test samples.

8.2 2013 Sampling for metallurgical testwork

Joseph A. Kantor of JAK Exploration Services, LLC supervised the collection of approximately

1,720 pounds (780 kg) of mineralized samples for the Longstreet project, compliant with NI 43-

101 QA/QC guidelines. The samples that were collected were used for the 2013 metallurgical

test program at McClelland Laboratories in Sparks, Nevada.

Eighteen large surface samples were collected from three historic test pits (six samples from

each). In addition, a total of 13 horizontal and 6 vertical channel samples were collected from

the underground Upper Adit. Geologically, the underground and surface samples represent

two distinct geological structural domains. One structural domain includes the Longstreet vein

(coincident with the Adit Fault) and its hanging wall. The second structural domain is the

footwall of the Longstreet vein. The current resource is hosted in both structural domains.

Underground sampling started about 180 feet in from the Upper Adit portal. Refer to Figure 8.1

for a diagram of adit sample locations. Horizontal samples are shown in Figure 8.1 as long

pencilled lines (along the northwest to southeast drift) and vertical samples shown as short

pencilled lines on east-west drift. Continuous 10-foot long horizontal channel samples were

______________________________________________________________________________________________


collected from 180 feet to 310 feet. A series of 6 vertical channel cuts, each approximately 6

vertical feet in length were collected every 10 feet along the vein in the westward drift.

A tungsten carbide-tipped saw was used to cut two parallel to sub-parallel two-inch to three-

inch deep slices in the adit wall. A sledge hammer and chisel were then used to take a

representative channel sample. Horizontal samples were labelled with the footage interval,

starting with 180 to 190 feet. From 180 to 270 feet, all samples were collected from the

western face of the adit. From 290 feet to 310 feet, sampling continued on the eastern face.

______________________________________________________________________________________________


Figure 8.1 Location of Underground Adit Samples

______________________________________________________________________________________________


The 13 continuous horizontal samples are each 10 feet long and the six vertical samples are

about six feet long from the back (top) of the drift (tunnel) to the floor of the drift.

Surface pit samples #1, #2 and #3 consisted of approximately 70%, 50% and 10% respectively,

from in-place pit walls with the remainder from loose blocks. These three pits were the source

of the original surface metallurgical samples used during the 1987 KCA testing. Based upon the

excavation outline in the pit walls, it appears that the original metallurgical samples consisted of

the silicified material with the high-clay content material avoided. For this 2013 bulk sampling

campaign, the pit #1 and pit #2 samples included high-clay content material in an amount about

equal to the bedrock exposure. Pit # 3 hosted very little clay-rich rock. Except for the clay-rich

samples, all samples collected were at least four inches to a maximum of about 10 inches in at

least one dimension. The mix of rocks collected at each pit was generally random and is

considered representative of the bedrock exposure. Refer to Figure 8.2 for location of surface

samples.

Figure 8.2 Approximate Location of the Three Surface Sampling Pits (shown in red)

______________________________________________________________________________________________


All of the bulk samples collected were either from surface exposures or at an approximate

maximum of 130 feet below the Upper Adit. No bulk samples were obtained from areas that

would be considered as the transitional or mixed oxide-sulphide zone.

8.3 Gold-Silver Mineralogy

An extensive search using Scanning Electron Microscopy/Energy Dispersive Spectroscopy

SEM/EDS indicates silver phases are present in both head and tail samples, however, gold was

not identified. Silver sulfide is the main silver phase and occurs as irregularly shaped inclusions

in quartz, pyrite and goethite pseudomorphs after pyrite. Cube-like grains are also seen in

quartz and likely represent pseudomorphs of acanthite after argentite. Grain size of the silver

sulfide is very fine with measurements that range from 0.5μm up to approximately 5μm. Silver

sulfide is also seen as thin rinds around pyrite and as small inclusions in jarosite. Much of the

jarosite in these samples analyzed by EDS contains low but detectable silver. The jarosite

contained in the samples is potassium jarosite, however, vague bright areas in large masses are

discernable using backscatter imaging. These areas are silver rich and likely represent

argentojarosite intimately mixed with the more abundant potassium variety. One small grain

having a chemistry of Hg, Br, Cl and Ag was identified as an inclusion in quartz with a

measurement just over 1 μm (one millionth of a meter). This phase may represent capgaronnite

or possibly iltisite. The primary reason for low silver recovery in this material appears to be due

to the very fine grained nature of the silver sulfide, which should leach easily if liberated or

exposed. In contrast, silver bearing jarosites tend to be refractory and are usually unaffected by

leaching.

8.3.1 Sulfide Mineralogy

Sulfides are present as a trace with pyrite as the main sulfide. Pyrite occurs as minute cubes and

drop-like grains that vary in size from <1μm up to approximately 20μm. Most grains are

unaltered but a small population wear thin goethite jackets. A trace of chalcopyrite is present

and shows no apparent decay.

8.3.2 Oxide Mineralogy

Both samples contain low amounts of iron oxide with hematite and goethite as the main iron

minerals. Hematite occurs as small rosettes, thin strings and small pockets. Goethite is generally

seen as euhedral pseudomorphs after pyrite. Yellow limonitic iron oxide is in the form of

irregularly shaped masses or intermixed with kaolinite. Secondary rutile forms small aggregates

and honey colored prisms in quartz.

8.4 2013 Metallurgical Test Program

The 2013 metallurgical test work program was conducted by McLelland Laboratories under the

direction of a QP metallurgical engineer contracted by Star Gold. The program included bottle

roll tests, column tests and comminution tests and mineralogical examination.

______________________________________________________________________________________________


8.4.1 Section Sample Assays

A total of 65 underground adit samples weighing 816 pounds (370kg) and three surface samples

weighing 904 pounds (410kg) were collected for metallurgical testing. Each of these samples

were crushed to 100% -2 inches (50mm) and assayed for gold and silver in duplicate. Assay

results are listed in Table 8.2. Samples were combined to generate surface and underground

composites, as well as a blended master composite. Triplicate direct assays were conducted on

each composite. Standard deviations between triplicate head assays were high, particularly for

the surface master composite. The agreement between the triplicate splits was not good,

however the average of the triplicate assays is close to what was expected, based on the

section assays. It was noted that the Quality Control samples all checked out as well, which

indicates that the assays are good and the gold occurrence in the potentially economic

mineralization is just a little “spotty”.

Table 8.2 Gold Head Assays and Head Grade Comparisons

Longstreet Composites

SMC, g/mt

UMC, g/mt

BMC, g/mt

Determination Au Ag

Au Ag

Au Ag

Direct Assay, Init. 0.21 17

0.70 67

0.57 40

Direct Assay, Dup. 0.67 34

0.82 63

0.66 41

Direct Assay, Trip. 0.37 21

1.09 53

0.77 50

Average 0.42 24

0.87 61

0.67 44

Std. Deviation 0.23 9

0.20 7

0.10 6

A total of twenty pieces of rock from both underground and surface were selected for

comminution testing. The remainder of the samples were separately stage crushed to 100% -2-

inches (-50mm). Each of the underground and surface samples were then blended to form a

master composite representing both the underground and surface samples. The blended

sample was then split to generate a third master composite. Samples were collected for bottle

roll tests. All composites were then further crushed to 80% -3/4 inch (19mm), blended, then

split into 75kg lots for column testing.

Selection sample assay results and detailed blending procedures are provided in the Appendix

to this report.

8.4.2 Bottle Roll Testing

A bottle roll test was conducted on each of the three composites at an 80% -10 Mesh (1.7mm)

feed size to determine lime requirements for column leach testing. Gold and silver recoveries

were similar for all three composites. Gold recoveries ranged from 80.6% to 81.9% and silver

recoveries ranged from 17.5% to 20.0%.

______________________________________________________________________________________________


Additional bottle roll tests, at a cyanide concentration of 1.0g NaCN/L were conducted on the

blended master composite at feed sizes of 100% -2 inches (50mm), 80% -3/4 inches (19mm) and

80% -1/4 inch (6.3mm) to determine sensitivity to feed size. The blended master composite

showed a moderate sensitivity to feed size with respect to gold and silver recovery. Recovery

was 18.4% higher for gold, and 13.9% higher for silver, at a feed size of 80% -1/16 inches

(1.7mm) than at a feed size of 100% -2 inches (50mm).

Silver recovery, for each bottle roll test conducted, was low. In order to investigate the cause of

the low silver recovery, three additional bottle roll tests were conducted on the blended master

composite to determine response to increased cyanide concentration (5.0g NaCN/L) at typical

heap leach (80% -3/4 inches, 80% -1/4 inches) and milled (80% -200 Mesh (75µm)) feed sizes.

Results showed that increasing the cyanide concentration did not significantly increase silver

recovery at heap leach feed sizes, however, silver recovery increased substantially when feed

was finely ground. Silver recovery was 60.6% from the bottle roll test conducted on 80% -200

mesh material. Gold recovery was also moderately higher when fine grinding was employed.

Mineralogical analysis of head and tail samples of the blended master composite confirm that

the primary reason for low silver recovery is due to the very fine grained nature of the silver

sulfide, which when exposed, is readily leachable. The silver leach rate at 200 mesh was

extremely fast. Silver recovery was complete within the first two hours, which suggests that the

silver mineralization is very fast leaching once liberated. In contrast, silver-bearing jarosites tend

to be refractory and are usually unaffected by leaching regardless of the grind size.

Summary results from bottle roll testing are given in Table 8.3. Detailed bottle roll test data including leach rate figures, are provided in the attached spreadsheet.

Table 8.3 Bottle Roll Test Results, 2013

Both gold and silver recoveries are slightly improved with increased crush size, the increase in recovery is more pronounced in the silver as compared to gold when a fine grind is applied. Figure 8.3 illustrates this. It is important to keep in mind that in order to reduce the particle size to 80 % passing 75 microns a conventional comminution circuit employing crushing and grinding would be required.

NaCN Au Ag

Feed Conc. Recovery, Calculated Head Recovery, Calculated Head

Composite Size g/L % Extracted Tail Head Assay % Extracted Tail Head Assay NaCN Cons. Lime Added

SMC 80%-1.7mm 1.0 80.6 0.25 0.06 0.31 0.42 20.0 5 20 25 24 0.08 2.1

UMC 80%-1.7mm 1.0 81.9 0.68 0.15 0.83 0.87 18.9 10 43 53 61 0.13 3.4

BMC 100%-50mm 1.0 62.9 0.44 0.26 0.70 0.67 3.6 2 54 56 44 0.07 1.3

BMC 80%-19mm 1.0 67.1 0.51 0.25 0.76 0.67 12.8 5 34 39 44 0.07 2.1

BMC 80%-6.3mm 1.0 77.9 0.53 0.15 0.68 0.67 13.6 6 38 44 44 <0.07 3.0

BMC 80%-1.7mm 1.0 81.3 0.52 0.12 0.64 0.67 17.5 7 33 40 44 0.13 2.5

BMC 80%-19mm 5.0 76.4 0.55 0.17 0.72 0.67 14.6 6 35 41 44 0.48 1.0

BMC 80%-6.3mm 5.0 77.6 0.45 0.13 0.58 0.67 14.0 6 37 43 44 0.67 1.0

BMC 80%-75µm 5.0 88.7 0.47 0.06 0.53 0.67 60.6 20 13 33 44 0.91 1.3

kg/mt ore

Table 1. - Summary Metallurgical Results, Bottle Roll Tests, Longstreet Mine Composites

Reagent RequirementsgAu/mt ore gAg/mt ore

______________________________________________________________________________________________


Figure 8.3 Crush Size Versus Metal Recovery

8.4.3 Column Leach Testing

Column leach test were conducted on each of the master composites, utilizing a feed size of

80% -3/4 inch (19 mm) in order to determine gold and silver recoveries, recovery rates and

reagent requirements under simulated heap leach conditions. Lime additions were based on

bottle roll tests. Test columns were sized at 15 cm diameter by 3 meters high using PVC piping

with material stacked in the leaching columns in a manner in which to minimize particle

segregation and compaction. Leaching was conducted by applying a cyanide solution of 1.0g

NACN/L over the charge at a feed rate of 12 Lph/m2 of column cross sectional area. After

leaching, fresh water rinsing was conducted to remove residual cyanide and to recover

dissolved gold and silver values.

Detail column leach tests data, including screen analysis of the feed and tails and drain down

rates can be found in the Appendix, identified as McLelland Report No. 3829 entitled Heap

Leach Cyanidation Testing Longstreet Project, dated April 6, 2014.

All three composites were leached for 190 days. Gold and silver extractions for the surface

master composite (SMC) reached 88.9 % and 20.0 %, respectively. Gold and silver extraction for

the underground master composites (UMC) was 84.6 % for gold and 15.4 % for silver. The

master blend composite (MBC) achieved gold and silver recoveries of 86.3 and 16.7 respectively.

Summary results from column leach testing are provided in Table 8.4. Detailed results, including

leach rate figures are provided in the Appendix.

______________________________________________________________________________________________


Table 8.4 Summary Metallurgical Test Results

Recovery results by size fraction for all three master composites indicates that finer crushing

would not substantially improve gold recovery. Gold recovery was similar throughout the

various size fractions with only a slightly elevated recovery in the finest size fraction (-75

microns). Silver recovery on the other hand would benefit from a finer particle size and would

require fine grinding in order to maximize recovery.

Overall metallurgical results indicate that the Longstreet master composites are readily

amenable to simulated heap leach treatment at 80 % -19 mm feed size. Gold recoveries for all

three composites were similar and ranged from 84.6 % to 88.9 % in 190 days of leaching and

rinsing. Silver recoveries were similar for all three samples, with recoveries ranging from 15.4 %

to 20.0%.

It is important to note that although the column tests were conducted over a period of 190 days,

gold extraction was essentially completed in the first 30 days of leaching. Silver leach rates, on

the other hand, were very slow and it is not expected that they would improve beyond the 190

day cycle.

Cyanide consumption rates were high and ranged from 1.56 to 1.93 kg NaCN/t of ore. This was

due in part to the long leach times. Cyanide consumption rates in a commercial operation are

typically much lower.

Figures 8.4, 8.5 and 8.6 diagramatically illustrate the leach rates and results for gold and silver.

Summary Metallurgical Results, Column Percolation Leach Tests, Longstreet Mine Composites,

80%-19mm Feed Size

Sample

I.D.

Test

No.

Leach/rinse

Time, days

mt/mt

ore

g Au/mt ore

Extracted

Average

Head

g Ag/mt ore

Extracted

Average

Head

NaCN

consumed

kg/mt ore

Lime

added

kg/mt ore

SMC P-1 153 4.8 0.32 0.38 5 24 1.45 1.7

UMC P-2 158 5.3 0.59 0.85 7 60 1.90 2.7

BMC P-3 158 5.2 0.63 0.68 8 45 1.78 2.0

______________________________________________________________________________________________


Figure 8.4 Surface Master composite leach kinetics

Figure 8.5 Underground master composite leach kinetics

0

10

20

30

40

50

60

70

80

90

100

0 30 60 90 120 150 180 210

Cu

mu

lati

ve

Rec

over

y,

% o

f T

ota

l

Leach Time, days

Gold and Silver Leach Rate Profiles, Column Leach Test,

Longstreet Mine, Surface Master Composite, 80% -19mm Feed Size

Au Ag

0

10

20

30

40

50

60

70

80

90

100

0 30 60 90 120 150 180 210

Cu

mu

lati

ve

Rec

over

y,

% o

f T

ota

l

Leach Time, days


Longstreet Mine, Underground Master Composite, 80% -19mm

Feed Size

Au Ag

______________________________________________________________________________________________


Figure 8.6 Master blend composite leach kinetics

8.4.4 Comminution Tests

Sample Preparation

A total of twenty competent pieces of rock were taken from the 22 samples for comminution

testing. Half of the 20 rock pieces were selected from the underground adit samples, and half

were taken from the surface samples. The rock pieces were combined and then submitted for

crusher work index and abrasion index testing.

No preparation was required for the crusher test sample. Pieces were natural rock and

fragments were used for the abrasion test. The abrasion test sample was crushed and screened

to extract a ¾ inch x ½ inch size fraction.

Crusher Work Index Test

The crusher work index test was conducted on natural rock pieces according to test protocol.

Sample CWi (kW-hr/st) CWi (kW-hr/mt)

Crusher Work Index 10.08 11.11

Abrasion Index Test

An abrasion index test was conducted on a -3/4 inch +1/2 inch fraction of the sample according

to test protocol yielding a Sample Abrasion Index of 0.2431.

0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

80.0

90.0

100.0

0 30 60 90 120 150 180 210

Cu

mu

lati

ve

Rec

over

y,

% o

d T

ota

l

Leach Time, days


Longstreet Mine, Blended Master Composite, 80% -19mm Feed Size

Au Ag

______________________________________________________________________________________________


8.4.5 Mineral Processing Conclusions:

AMPL concludes the following from the results of metallurgical testwork and analysis:

Cyanide consumption was high and ranged from 1.56 to 1.93 kg/t, in part due to the long leach time (190 days) used in the column test. Typical cyanide consumption for a heap leach operation would be 25% to 40% of the consumption observed in column leach tests.

Cyanide consumption rates were significantly lower for the bottle roll tests.

Due to the coarse crush size and low amount of generated fines it can be assumed the ore will respond well to permeability testing during the next phase of column tests, thus eliminating the need for agglomeration. However if a finer crush size is tested then it may require incorporating cement in the agglomeration mix.

Column tests indicate that gold dissolution is rapid with very little additional recovery achieved after 30 days of leaching. Silver leach kinetics were slow and continued to increase slightly even after 120 days of leaching.

Results indicate that gold recovery is not particularly sensitive to feed size, given a sufficient leach cycle time.

Each of the three master composite samples exhibited amenability to simulated heap leaching at a particle size of 80% minus ¾-inch (19mm). Gold recoveries in this size fraction ranged from 84.6% to 88.9%.

Column test silver recoveries were low, ranging from 15.4% to 20.0%.

The crusher work index for the ore indicates it to be of low hardness and slightly abrasive.

Increasing cyanide consumption from 1 g/L to 5 g/L in bottle roll tests had little impact on both gold and silver recovery at varying crush sizes.

8.4.6 Mineral Processing Recommendations

Further column test work on the oxide adit material should be performed in order to test the variability of the deposit.

Further column leach tests should be conducted using finer material in conjunction with high pressure rolls i.e. P80 ¼-inch (6.3mm) in order to maximize silver recovery.

As the leach kinetics for gold are fairly rapid and the silver recovery did not increase dramatically after 190 days of leaching, it is recommended to reduce the column leach time to 60 days for the next phase of the test work.

Further column tests should be carried out using site water as opposed to laboratory tap water in order to determine the effects of site water on leach kinetics.

To investigate improved silver recovery on the master blend composite ore, a HPGR (high pressure grinding rolls) evaluation should be considered as HPGR’s lead to the formation of micro cracks in the ore which may improve silver leaching kinetics. This would require a Static Pressure Test (SPT) to be performed.

______________________________________________________________________________________________


Load/permeability tests are recommended on column leach residue samples to confirm permeability under compressive loading.

8.5 Process Engineering and Design

The process layout and equipment selected for the Longstreet heap leach study is based on the

2013 metallurgical test program which was limited to several bottle roll tests, percolation tests,

hardness and abrasion index determinations and column tests conducted on three composite

sample. AMPL also referenced previous metallurgical results from the 1988 pre-feasibility study.

The process plant design, which includes the leach pad, ADR plant, electro-winning circuit and

refinery is based on a nominal five year mine life. In order to maximize project efficiencies and

minimize capital and operating costs, a plant utilizing modular components should be

considered.

The proposed crushing facility and leach pad stacking would be operated by an independent

contractor with the crushing plant consisting of a two-stage modular design and haul trucks

used to stack material on the leach pad. For the sole purpose of this conceptual study, the heap

leach pad and processing plant for the Longstreet Project is designed to process 2,800 t/d of

low grade gold and higher grade silver run-of-mine (ROM) material. Both the crushing and

stacking areas will operate 16 hours per day, seven days per week at 90% availability.

The adsorption-desorption-refining (ADR) facility should also be of modular design in order to

minimize capital cost and reduce the construction schedule. The life of mine recovery rate for

gold is estimated at 86% while silver recovery is estimated at 15%. The metals recovery plant

(ADR) facility is designed to treat a solution flow rate of 770 gpm (175 m3/h) of pregnant leach

solution which will produce approximately 18,700 ounces of gold and 78,800 ounces of silver

per year. The ADR plant will operate on a 24 hour per day basis, seven days per week at 90%

availability.

A summary of the design criteria for the heap leach and ADR facilities is presented in Table 8.5.

Table 8.5 Process Design Criteria

Design Criteria Design Parameters

Ore to leach pad 1,000,000 tpy

Maximum rock size to crusher Minus 24 inch

Nominal crushing rate 210 t/h

Design crushing rate 240 t/h

Crusher Work Index 10.08 kWh/short ton

Abrasion Index 0.2431

Ore moisture content 3%

Ore moisture content during leaching 12%

Final crush size to leach pad 80% -3/4 inch

______________________________________________________________________________________________


Annual operating days 365

Crusher availability 90%

Crushing – hours per day 16 h/d

Stacking – hours per day 24 h/d

ADR plant – operating hours per day 24 h/d

AR plant availability 90%

Carbon tons per column 4.4

Number of CIC columns 5

Tons of carbon transferred per day 2.2

Heap configuration – starter pad 3 lifts – 26 feet per lift

Starter pad – Number of cells per year 4

Tons per cell 246,575

Leach cycle – primary and secondary cells 180 days

Solution flow to ADR plant 776 USGPM

Solution flow to primary leach cell 810 USGPM

Solution flow to secondary leach cell 810 USGPM

PLS pond capacity – live volume 972,150 US gal

BLS pond capacity – live volume 2,138,737 US gal

Solution application rate to leach pad 0.27 US gal/h/ft2

Gold recovery, estimated 86%

Silver recovery, estimated 15%

A simplified block diagram of the process is shown in Figure 8.7. The heap leach plant will

employ two stage crushing with a jaw and a cone crusher. Crushed product will be trucked to

the heap leach where cyanide solution will be added. Pregnant leach solution will percolate

through the heap and eventually be pumped to the ADR plant which will consist of a series of

carbon contactors, elution column, acid column and rotary kiln. The recovery plant will house

an electro-winning cell along with a bullion furnace.

______________________________________________________________________________________________


Figure 8.7 Flowsheet block diagram

A simplified overall flow sheet of the process is illustrated in Figure 8.8.

Flowsheet Block Diagram

Crushed Ore

P100= 500 mm

Oversize material

Undersize material

Undersize material Heap Leach Pad Feed

F100 < 19 mm

Lime Addition

Cyanide Addition

Barren

Solution

Loaded Carbon Barren Carbon

Pregnant Solution

Barren Solution

High Grade Sludge

Dore Bars

Crusher Hopper Stationary

Grizzly

Pregnant SolutionPond

Heap Leach Pad

Jaw CrusherPrimary

Double Deck Screen

Cone Crusher Secondary

ElectrowinningCircuit

Carbon Elution Circuit

Carbon Regeneration

Circuit

Barren SolutionPond

Refinery Includes:Calcine Dryer &

Bullion Furnace

CIC Circuit

Grizzly Feeder

______________________________________________________________________________________________


Figure 8.8 Flow sheet diagram

Run of mine ore

Oversize + 600 mm

Crusher Feed Rock Breaker

Hooper

Grizzly Feeder Primary Jaw Crusher

Magnet

Double Deck Screen

Transfer Conveyor

Screen Feed Conveyor

Lime Addition Secondary Cone Crusher

Screen Product Conveyor Crusher Discharge Conveyor

Pregnant Solution Pond Pumps

Barren Solution Pond Pregnant Solution Pond

To Atmosphere

Storm Pond

Loaded Carbon Loader Carbon Recovery Screen Furnace and Dryer off

Gas Scrubber

Carbon Elution Column Calcine Oven

Acid Wash Calumn Barren Solution Tank

Pregnant Solution

Ellectrowinning Cell High Grade

Rotary Kiln Slugde

Barren Carbon Dore Bar

CIC Tank # 1 Bullion Furnace

CIC Tank # 2 Carbon Quench Tank

Regenerated & Fresh Carbon

Fresh Carbon

CIC Tank # 3

Carbon Sizing Screen

CIC Tank # 4

CIC Tank # 5

Carbon Safety Screen

Barren Solution from Electrowinning Cell

Cyanide Solution

Barren Solution Tank

______________________________________________________________________________________________


8.6 Process Description

The process of winning gold from the mined ore involves a number of individual activities,

described in detail below.

8.6.1 Crushing

It is recommended that the crushing plant be a two-stage closed circuit crushing

facility. The crushing facility should be specified to receive ore at 24-inch (600 mm) top

size and crush it to 80% passing ¾ inch (19 mm). The design crusher feed rate is based on

processing an average of 1,000,000 tons per annum, operating 16 hours per day, 365 day per

year, for an operating availability of 90%, which equates to 210 tons per operating hour.

The remaining eight hours of the day will be used for maintenance.

ROM material will be dumped by haulage trucks into a hopper equipped with a grizzly and

rockbreaker (for oversize) or placed in a stockpile situated close to the hopper for reclaim to the

hopper by front end loader.

8.6.2 Double Deck Screen

Below the hopper a vibrating feeder will feed a double deck screen located above a jaw

crusher. Double deck screens are strongly recommended for the crushing circuit due too

the high recirculating load and the wide size distribution of material fed to the screen.

The screen will divert finer ore away from the jaw crusher to the final product conveyor

belt, thus improving crusher efficiency. Screen oversize material will feed a short-head

cone crusher for further size reduction and then be recirculated back to the double deck

screen.

The bottom deck screen undersize is the final crushed product, which is conveyed to the

crushed ore stockpile. The bottom deck is sized to produce a heap leach feed size which is

specified as 80 percent of the ore being less than ¾ inches in size.

8.6.3 Lime Addition

Lime will be stored in a silo adjacent to the belt conveyor where it will be added to the crusher

plant output conveyor, where it will be thoroughly mixed with the ore. Lime is used to

agglomerate the ore and for pH control in the heap leach.

8.6.4 Heap Leach Pad Stacking

The ore is stacked on the pad, using the open pit haul trucks, in a number of lifts in a

pyramid type layout. Each lift will have a 26 foot (8 m) setback all around the previous lift.

This setback allows a safety berm to catch material that may slough off the lifts placed

above and minimizes the risk of spills of cyanide-containing solutions and cyanide

soaked ore or spent ore from the containment system. It is estimated that the ultimate

heap leach pad will have a foot print of approximately 1,600 feet by 790 feet (490 m by

240 m) and measure approximately 160 feet (48 meters) in height. Pad dimensions need to be

______________________________________________________________________________________________


verified for the next phase of the project.

The starter pad will take two years to complete and consist of approximately 2 million tons of

ore. Each successive lift will be placed on top of the previous lift and will be setback from

the crest inside edges of toe and perimeter berms to provide corridors for solution

application pipelines and access. This will provide the second lift and all future lifts with a safe

access for heavy equipment while providing extra room in case there is any slumping of the lifts.

Initially crushed waste rock would be placed onto the pad to provide a protective layer of

material on the pad liner for equipment to operate on, eliminate expansion and contraction of

the synthetic liner and prevent damage that may occur due to weather conditions such as

sun damage or ripping by winds.

Perforated pipe will be placed on the top of the crushed protective layer to aid in the flow of

pregnant solution from underneath the leach pad to the solution collection ditches. Heap

leach ore is then stacked onto the pad.

Stacking will occur in cells which are approximately 270 feet (82 m) in width. Each cell is 790

feet (240 m) in length and represents approximately three months of stacking, assuming 26 feet

(8 m) high lifts. As the stacking of the ore retreats from the stacked face and the entire

length of the pad has been stacked, ore is again transported to the far end of the leach

pad and a second cell or strip of ore is stacked adjacent to the completed cell.

8.8.3 Leachate Distribution & Collection

A barren leachate solution distribution line will run along the side of the leach pad. A

series of headers with valves will run from the barren line up onto the leach pad. The drip

emitters (apply leachate to the ore) will then be connected to the headers (in each

direction) and extended across the ore to distribute barren solution over the entire area

of the leach pad, for leaching of the precious metals from the ore. Blocks of ore are stacked

with lengths equal to the distance between headers on the barren solution distribution line.

As soon as a block of ore is stacked, the ore will be placed under leach. Getting ore under

leach as quickly as possible is key to maintaining production in a heap leaching operation.

As each cell is stacked the headers are extended and fresh ore is placed under leach.

Drip emitters are well suited for dry climates as they reduce water losses by evaporation.

Barren solution lost to evaporation is replenished with makeup water containing cyanide.

Minimizing water consumption is an important aspect of this project. Drip emitters have

the added advantages of providing good solution flow rate control, excellent solution

penetration of the heap and, generally, more even solution distribution than other

methods of solution application.

The disadvantage of drip emitters is the susceptibility to blockage by suspended particulates or

possible scale formation. Anti-scalant is added to prevent or minimize scale formation. If scale

formation from evaporation becomes a problem, the drip emitters can be buried below

______________________________________________________________________________________________


the surface of the stacked ore.

Once ore has been under leach for the assumed 180 day leach cycle, it will be removed from

leach as additional new ore is placed under leach. In this manner, a constant volume of

ore will be under leach at one time and the optimum solution application rate can be

maintained. This stacking and piping sequence is continued until the entire leach pad is

covered with the first lift of ore. A similar stacking and piping sequence will be followed until

the entire pad reaches its ultimate design height.

8.6.5 Solution Ponds

A pregnant solution pond will be constructed near the lowest point of the pad to store leachate

solution containing gold and silver and storm runoff flows from the pad. The pond will have a

bottom corner sump and a leak detection system between the geomembranes and connect to a

corner leak detection sump and well system. Solution from the pregnant solution pond will be

pumped to the ADR by two pregnant solution pumps, one operating and one spare. Total pond

capacity is estimated at approximately 1.3 million USG (5000 m3) of solution.

A barren solution pond will also be constructed near the lowest point of the pad to store

leachate solution and storm runoff flows from the pad. The pond will have a bottom corner

sump and a leak detection system similar to the pregnant solution pond. Solution from the

barren solution pond will be pumped to the heap leach. Total pond capacity is estimated to be

approximately 2.6 million USG (10,000 m3) of solution.

An emergency pond will be constructed adjacent to the pregnant solution pond, to facilitate a

major storm event emergency. The storm pond also increases the capacity of the stored

solution within the pad area. During storm conditions, all storm water and heap liquor is

retained in a closed circuit thus preventing spillage and subsequent solution losses to the

environment. This pond would not be used under normal operating conditions.

8.7 Adsorption, Desorption and Refining (ADR) Facility The winning of gold from the pregnant solution occurs in the ADR plant. The steps in the gold-winning

process are:

8.7.1 Adsorption circuit

The carbon adsorption circuit is based on a five-stage, upflow CIC system, Solution from the

pregnant solution pond is fed to the ADR plant. The carbon columns are designed for 100%

carbon bed expansion. The carbon adsorption circuit consists of a series of five cascading

carbon columns. Each column is designed to contain four tonnes of activated carbon. Solution

enters the circuit at the first carbon column and flows counter-current to the flow of carbon.

Solution overflows the final column onto the stationary carbon safety screen to catch any

entrained carbon.

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The counter current flow allows the carbon in the first adsorption tank to reach optimum metal

loading. Design loadings are 3,500 g/t gold and silver, but actual loadings will be a function of

the solution grades reporting to the ADR circuit. Two tonnes of loaded carbon are advanced per

day from the first CIC tank to the acid wash circuit.

The barren solution that discharges from the final carbon column drains to the carbon column

surge tank via a carbon safety screen. From this tank, barren solution is pumped back to the

barren solution pond of the leach pad.

Loaded carbon is passed over a loaded carbon recovery screen prior to entering the acid wash

tank, allowing the solution to return to the CIC circuit. Fresh and regenerated carbon will be

introduced into the CIC circuit via the last carbon column at the same rate the loaded carbon is

removed, thus maintaining a constant carbon inventory.

8.7.2 Carbon acid washing

Loaded carbon is passed over the loaded carbon recovery screen into the acid wash tank where

any scale buildup on the surface of the carbon is removed. A 3% w/w hydrochloric acid (“HCI”)

solution is circulated through the vessel at a rate of two bed volumes per hour for 120 minutes.

Upon completion of the acid rinse, the carbon is soaked in the HCI solution for an additional 60

minutes. After soaking, the spent acid and carbon are neutralized with sodium hydroxide

(NaOH) solution. The spent acid solution is sent to the CIC circuit, and the acid washed carbon is

pumped to the carbon elution column.

8.7.3 Desorption circuit

A barren strip solution (used to remove gold and silver from the carbon) is heated to

approximately 80°C in a diesel-fired single-pass indirect solution heater. Enough sodium

cyanide and sodium hydroxide is added to the heated barren solution so that a 1% by weight

sodium cyanide and 1% by weight sodium hydroxide solution is obtained. The barren strip

solution temperature is then increased to approximately 115°C, prior to entering an elution

column where metals stripping is completed. In order to enhance the desorption of the gold

and silver from the loaded carbon, the loaded carbon is also soaked 30 minutes in the 1%

sodium cyanide and 1% sodium hydroxide solution. After soaking, the carbon and solution are

sent to the elution column. During the metals stripping process, 40-bed volumes are passed

through the bed of loaded carbon in the elution column. During the stripping cycle, the loaded

strip solution is continuously circulated from the elution column to the electro-winning circuit.

It will take approximately 21.5 hours to complete one strip cycle which includes carbon

transfers, acid washing, carbon stripping and electro-winning. The final stage of elution

will include an unheated wash water cycle that will displace the last bed volume of

pregnant strip solution and, in turn, cool off the barren carbon.

The pregnant strip solution is sent to electro-winning cells for further processing. The cooled

carbon is transferred to the horizontal rotary kiln for thermal regeneration.

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8.7.4 Carbon thermal regeneration

The carbon regeneration circuit is sized to handle a carbon transfer rate of 2 tonnes per day,

with all barren carbon thermally regenerated prior to reuse in the CIC circuit. Carbon is

dewatered with a static sieve bend screen before entering the drying kiln at a rate of

approximately 330 lbs (150 kg) per hour. The horizontal kiln is designed to have a retention

time of 20 minutes and a bed temperature of 1560°F (850°C). The majority of organic

compounds fouling the barren carbon will be removed in this process. When the

regenerated carbon exits the kiln, it is immediately deposited below water in the carbon

quench tank.

Regenerated and fresh carbon are transferred back to the CIC circuit. All carbon is screened

prior to reuse in the CIC circuit in a static sieve bend screen located over top of the No. 5

CIC tank. Under sized barren carbon is collected in a barrel for recycling.

8.7.5 Refining

Pregnant strip solution, stored in the loaded solution tank, is pumped to a single

electro-winning cell. Each electro-winning cell has eleven 316 stainless steel wool cathodes

and twelve 304 stainless steel sheets to be used as the anodes. Gold and silver are plated onto

the steel wool cathodes in the electro-winning cell.

The cathodes are removed periodically from the electro-winning cells and the gold and

silver sludge is washed off using a high pressure spray. Sludge collected from washing the

cathodes and from the bottom of the electro-winning cells is first passed through a plate and

frame filter press in order to remove excess water and then dried in an oven.

The dried sludge is mixed with flux and charged into a diesel fired melting furnace for smelting.

The dore is poured out of the tilting furnace and collected in four cascade style molds. The slag

exits the melting furnace first and is later displaced by the heavier gold and silver dore. The

slag ultimately collects in the slag pot which is placed below the cascading molds.

The mineralogy report indicated an absence of mercury in the sampled material. Therefore a

mercury retort furnace is not required.

Electro-winning tails solution is returned to the barren strip solution tank with barren strip

solution is periodically bled from the barren strip solution tank and replaced with fresh

solution. This controls impurity levels in the barren strip solution. The barren strip solution

is returned to the carbon columns circuit.

8.7.6 Water services

Raw water will be pumped from wells to the water tank prior to distribution throughout the plant. Potable water will be sourced from the RO plant. The total raw water consumption will depend on seasonal evaporation rates which can range from 4% to 10%.

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8.7.7 Reagents

All applicable safety considerations will be made, including separation of acids and cyanide,

provision of safety showers and eye wash stations, and designated sump pumps.

Hydrochloric acid will be delivered to site in 50-gallon (200L) drums and diluted with water in a

tank. Sodium hydroxide will be delivered to site in 50 pound bags and mixed with raw water in

a tank. Cyanide will be delivered to site as solid briquettes of sodium cyanide in 1 tonne bulk

bags. Cyanide will be stored in the dry reagent storage area prior to mixing with water in a tank

to obtain a 20% w/w solution.

Activated carbon will be delivered in 500kg bulk bags and will be emptied into the carbon

quench tank to be mixed with raw water and newly regenerated carbon. Hydrated lime will be

delivered to site in 20 tonne trucks and transferred to a lime silo for storage.

8.7.8 Assay laboratory

It is assumed that all exploration and process plant samples will be sent to an external

laboratory for analysis.

8.8 ADR plant manpower

The process plant will require 15 people including maintenance personnel. Contractors are not included in the total man power requirement.

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9 INFRASTRUCTURE

The Longstreet Gold Project lies in a relatively remote region of Nevada with sparse human

population and few towns, highways or power lines.

9.1 Site Access

The project site has a reasonable gravel road access adequate for an exploration project but

will have to be upgraded if the project advances to production. A paved county road runs E-

W approximately 27 miles south of the Project, connecting the site to the nearest town of size,

Tonopah, which lies 48 miles to the southwest.

Most of the Longstreet Exploration Project is located within the Georges Canyon Inventoried

Roadless Area (IRA), which topic is discussed in Section 11, below.

9.2 Power and Power Distribution

At present, there is no electric power, telephone or internet service on or close to the site.

Therefore, required electrical power will have to be generated with diesel-powered

generators. Approximately 700 kW installed power is required for the proposed heap leach

operation. A single 1.0 MW heavy fuel oil (“HFO”) driven generator would be able to supply the

heap leach and ADR plant. It is assumed that the power for the crushing plant will be supplied

by the contractor. Electricity will be distributed across the complete site via 6.6 kV overhead

power lines.

Power is not distributed to the water well intake pumps due to the distance from the power

plant. Mobile generators will be used for powering this facility.

9.3 Site roads

An allowance for site roads connecting surface support facilities at the open pit and heap leach

site have been included. The roads would be built from open pit waste rock, gravel top covered

and wide enough to accommodate 2 way vehicle traffic.

A haul road from the open pit top the heap leach pad would be constructed from open pit

waste rock with a gravel cover and wide enough to accommodate 2 way haul truck traffic.

9.4 Surface support buildings

Office space for the limited technical, surface support and administrative staff of the company

would be housed in several office trailers placed on site and provided with electricity, water and

sewage services. Conference room and washroom facilities would also be provided for the

office space.

A prefabricated building or converted shipping containers with concrete floors would be

equipped as a mine equipment maintenance shop and warehouse for servicing the project.

An explosives magazine for powder and detonators will be constructed at acceptable distance

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from the mining operations and other surface buildings and facilities.

All entry and exit from the property would be via a security trailer located bythe office complex.

It would house an area with turnstiles, a room for searching people to minimize theft and a first

aid room.

9.5 Other services

Telephone and internet communication infrastructure will have to be constructed and utilize

satellite communications systems. The site will be provided with computer servers and desktop

or laptop computers.

A fuel storage area equipped with diesel tanks and storage for oils will be constructed near to

the open pit.

Garbage will be hauled by contractor to the nearest licenced disposal site.

For this study water supplies have assumed to be from wells. The area is known to contain

springs and water at depth. Project water requirements need to be estimated and the source

of the required water determined, as this will be critical to project advancement.

Water management will include collection ditches and ponds and a water treatment plant.

Sewage will be processed in a septic and filtration system.

9.6 Area support services

Tonopah exhibits some support infrastructure for an open pit mining operation, including a

local workforce, some support contractors, shipping facilities, etc. Other required services

can be sourced within the region.

9.7 General and administrative

General and administrative (G&A) costs are those primarily associated with the general

management and administration of the project. G&A is associated with surface facilities and

personnel not included under the mining, product preparation or maintenance groups and in

addition to the surface department comprise of: administration; procurement; human

resources; and security.

9.7.1 Administration

Administration comprises senior and general management, accounting, third party

environmental support and information technology functions. In addition to employee salaries

and benefits, other components include employee relocation, travel expenses for business

away from the property, insurance (property and business interruption), permits and licences,

fees for mining rights, professional fees, and operating surface vehicles for the personnel.

Accounting functions include payroll, accounts payable, accounts receivable, budgeting,

forecasting and other corporate cost accounting.

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Information technology comprises all components associated with operating and maintaining

the telephone, computer network, internet, fax and radio systems for the mine site.

Allowances for long distance telephone charges are also included.

Environmental costs are associated with monitoring of the mine’s environmental performance

and reclamation work.

9.7.2 Procurement

Procurement encompasses all functions associated with on and off site procurement of

materials and supplies; warehousing and inventorying; transportation from point of origin to

site and other associated support services. Actual freight costs for items required by the mine,

processing plant and maintenance departments are included in those department’s costs.

The main cost components are comprised of employee salaries and benefits and warehouse

supplies (such as personal protective equipment). Also included is small equipment (pallet

lifters, forklifts, etc.) and parts used for warehousing, purchasing and logistics. Surface support

includes loading and unloading of trailers and shipping containers, movement of materials on

site and maintenance of the warehouse and associated facilities.

9.7.3 Human Resources

Human resources encompass all functions associated with personnel, union relations, health

and safety, training and community relations. Personnel and industrial relations costs include

salaries and benefits for employees to recruit required personnel, manage Company salary and

benefits policies, manage hourly employees and oversee the Company’s policies and

procedures. Health and safety includes salaries, benefits, on-site first aid personnel, first aid

supplies and vehicles required by this group.

Community relations costs include funds to aid in supporting local community efforts and

facilities.

9.7.4 Security

Mine site security is provided on a contract basis by a third party security firm. Security

surveillance equipment will be provided to the security firm by the mine. Other minor security

equipment for the security personnel (such as metal detectors, etc.) would be provided by the

contractor. Thesecurity facility would be constructed at the entrance to the mining areas and

by the office complex,to prevent inadvertent access to the mine site. All personal vehicles will

be parked at security and transportation by bus will be provided to the mine site for the work

force.

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9.7.5 Manpower

The G&A manpower required for the mine after commercial production starts is estimated to

be 13 employees with the cost structure based on expected salaries paid in the US mining

industry. The G&A manpower is presented in Table 9.1.

Table 9.1 G&A Personnel Complement

Position Complement

Mine manager 1

Senior engineer 1

Accountant 1

Engineering/Geology technicians 2

Purchasing/warehouse manager 1

Environmental co-ordinator 1

Medical contract 1

Security guard 4

Site services 1

Total 13

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10 HYDROLOGY

10.1 Water Sources

Due to the limited information on the available water within the Monitor Range and at the

Longstreet Project site, a hydrogeologic evaluation of water availability from sources both on

and off the Project site is strongly recommended.

The hydrogeologic evaluation would consist of three sections:

1) Onsite hydrogeologic evaluation

2) Offsite hydrogeologic evaluation

3) Stone Cabin Valley hydrogeologic evaluation

The proposed outline of the hydrogeologic evaluation follows:

1) The onsite hydrogeologic evaluation will include at a minimum:

a. Perform detailed structural mapping of fracture, fault and/or joint system(s)

associated with each of the lithologic units described in Section 4 of this

document.

b. The rock quality designation (RQD) determined of the diamond cores from

previous mineral investigations to measure the degree of jointing or fracture in

the various lithologies.

c. Utilize structural mapping to create a hydrogeologic model of the mine site to

predict potential locations for groundwater.

d. Install test wells to perform pumping tests to evaluate the volume of available

water at the mine site; and predict the available volume of water for sustained

mine and plant operation.

e. Sampling the groundwater to establish a water quality baseline, and confirming

the quality for mine and plant operation.

2) The offsite hydrogeologic evaluation will include at a minimum:

a. Geologic/hydrogeologic mapping to evaluate possible production well locations

within 1.5 miles of the Longstreet Mine site.

b. Evaluate the Side Hill Spring area; including if the water source is the

unconsolidated alluvium or the bedrock.

c. Install a test well at Side Hill Spring to perform a pumping test to evaluate what

effects pumping the spring would have.

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d. Sample the spring water and the groundwater to establish a water quality

baseline, and confirm the suitability for mine and plant operations.

e. If, in the course of these evaluations, other potential water sources are

identified, the same evaluations should be applied.

3) The Stone Cabin Valley hydrogeologic evaluation will include at a minimum:

a. Identify potential well(s) locations.

b. Permit requirements.

c. Test well installation and pumping test to ascertain the number of wells

necessary to provide water for the mine and plant operation.

d. Detailed cost estimate of well installation, operation, and pipeline installation.

The estimated cost to perform the necessary hydrogeologic evaluation ranges from US$200,000

to $325,000. It is dependent upon the number of wells installed, tested, and sampled.

Current information indicates that Stone Cabin Valley is the only known source available for

long term groundwater use. Due to the Valley’s distance from the mine, it may be the most

expensive option. The location of the well(s) is an important factor to evaluating costs; it is also

critical to secure long term water production that is not affected by variations in annual

recharge. The location for the first test/production well(s) in Stone Cabin Valley is the centre of

the valley; this is approximately 5.5 miles from the mine site. The hydrogeologic study and the

testing may indicate that a production well is feasible approximately 1 to 2 miles closer to the

mine site: An obvious reduction in pipeline construction costs, and maintenance. The estimate

for the installation of a pump station (2), well, pump, electrical supply, storage tank and pipeline

ranges from $1.4 m to $2.65 m depending on the distance from the mine (3 miles to 5.5 miles).

10.2 Water Usage

Water usage at a mine comprises mine operations (drilling, dust control, core shack, equipment

cleaning, etc.) use, plant operations use, and human consumption (drinking, showers and

toilets) use. Currently no demand estimates have been provided for mine operations or human

consumption use; the estimate provided for plant operations use follows: The water

demand for the heap leach plant operation is estimated to range from 20 to 45 m3/hour

(includes reagent requirements and make up water). The operation and maintenance costs for

pumping and transporting water include energy, operator and equipment from Stone Cabin

valley is estimated to be $230/day or a range of $0.21 to $0.48/m3.

Addition information is required for a complete water demand costs.

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10.3 Dewatering

The available information indicates that mine water management and dewatering does not

appear to be an issue at the Project site.

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11 ENVIRONMENTAL AND PERMITTING

Star Gold has staked and maintains 75 unpatented mineral exploration claims on United States

Forest Service (USFS) and Bureau of Land Management (BLM) lands. The Company had an

active Plan of Operations for its recent exploration work which included core drilling, surface

prospecting and gathering of bulk samples. The company is in the process of applying for a new

Plan of Operations for future exploration activity. Permits have yet to be applied to facilitate

full mining operations.

Past engineering work had proposed to locate the required leach pads on claims in a small

canyon adjacent or in close proximity to three potential ore zones. These claims are on USFS

lands. It has been noted that permitting on USFS lands is generally considered more difficult

versus BLM lands. Additionally water and pregnant solution management is more difficult in

the canyon versus flat terrain. To mitigate these potential hurdles, Star Gold staked a series of

claims on adjacent BLM lands that are flat and could be used for the leach pads if necessary.

This provides Star Gold with two possible options for leach pad location and we have assumed

this operating scenario in discussing the permitting process for the Longstreet Project.

11.1 Permitting Process

The Star Gold project is of modest size and the area where it is located is largely undisturbed or

has naturally reclaimed itself from past man-made disturbance. There are no pre-existing

environmental issues or liabilities on the site.

To assist the permitting process for mining and exploration activities in Nevada, a

Memorandum of Understanding (MOU) exists between the Nevada Division of Environmental

Protection (NDEP), the USDA Forest Service (USFS) and the US Bureau of Land Management

(BLM). It has been in place since 2008 and helps to coordinate the responsibilities of the

Agencies pertaining to the administration and reclamation of lands disturbed by exploration or

mining operations. Although this agreement expired in November 2013, AMPL has investigated

and been assured by the Bureau Chief of Mining Regulation and Reclamation of NDEP that the

various agencies are working on a renewal of this agreement and that any changes are expected

to be minor.

The following permits will be required from the USFS, BLM and NDEP for the mine to go into

production:

11.1.1 US Forest Service

Approval of a Plan of Operations;

Approval for upgrading access roads;

Approval of a reclamation plan for USFS lands with notice to NDEP (the

reclamation plan is part of the Plan of Operations);

Approval of a reclamation cost estimate for USFS lands for bonding purposes;

and

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An Environmental Impact Statement (EIS) (triggered by a request for the above

approvals). It is presumed that the BLM would be a Cooperating Agency or a

joint Lead Agency.

11.1.2 Bureau of Land Management

Approval of a Plan of Operations;

Approval for upgrading existing roads or granting rights-of way for new roads;

Approval of a reclamation plan (part of Plan of Operations) in a format that has

been developed jointly with NDEP;

An Environmental Impact Statement (EIS) (triggered by a request for the above

approvals). The EIS should be prepared in cooperation with the USFS as noted

above;

Approval of a reclamation cost estimate for bond purposes. (The cost estimate is

separately reviewed by NDEP); and

If a single bond is to be issued, the MOU noted above states that “…an interagency

agreement may be executed as necessary.”

11.1.3 Nevada Division of Environmental Protection (and other Agencies as noted)

Water Pollution Control Permit. This is a major permit in Nevada, required

whether or not there is any water discharge contemplated. Much of the

information required for Federal EIS purposes will serve as input for the

application for this permit. Analysis of the acid generating potential of all types of

rock to be disturbed is an important part of this permit. Also required for the

permit application are descriptions of the geological and hydrogeological

conditions, proposed operating plans, proposed monitoring plans, detailed

descriptions of leach pads and ponds, etc.;

Reclamation Permit. This permit application must utilize guidelines prepared by

NDEP and BLM (the USFS has its own guidelines). Cost estimates for carrying out

the plan by a contractor will be used to determine bond amounts. As noted

above, this is usually done in conjunction with the BLM and USFS. Bonding must

be obtained before construction can begin;

Storm Water Permit. This permit is a general permit requiring only application to

obtain coverage, but requires preparation of a Storm Water Pollution Prevention

Plan;

Air Quality Operating Permit. As pertaining to the Longstreet Gold Project, this

permit covers emissions from diesel generators, rock crushing and mining

operations;

A permit to appropriate water must be obtained from the Nevada Division of

Water Resources; and

An Industrial Artificial Pond Permit must be obtained from the Nevada

Department of Wildlife.

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11.1.4 Other Permits

There are several other permits which would be expected to be issued rather routinely with

minimal input from the applicant, as opposed to the above-listed permits, most of which will

require significant scientific and engineering input.

Other required permits include:

Permits to store explosives and cyanide;

Permits to treat sanitary waste and dispose of plant and office trash on-site;

Permit if a drinking water system is to be installed;

County permits such as business license and building permits; and

Registration with various agencies.

Once mining commences, a Toxic Release Inventory must be filed annually with the US EPA and

the Nevada State Emergency Response Commission.

11.2 Timing of Approvals

Based on some recent permitting in Nevada the time required to secure permits prior to the

construction of facilities and mine pre-stripping is estimated as two-to-four years from the

beginning of the environmental baseline studies. This assumes no significant objections are

raised by members of the public including indigenous peoples, environmental groups or other

government agencies (note that the US EPA conducts a review of all environmental impact

statements). It also assumes that no significant issues arise with respect to cultural resources or

endangered or sensitive species. (Note that no endangered or threatened species or cultural

resource issues were identified in site surveys conducted to receive USFS approval for the

exploration drilling program, although some restrictions were imposed due to the possibility of

the presence of sage-grouse. The sage-grouse issue is discussed further on in this section.)

11.3 Inventoried Roadless Area

According to the USFS Decision Memo of August, 2011, most of the Longstreet Exploration

Project is located within the Georges Canyon Inventoried Roadless Area (IRA). While noting that

the project area is open to entry under the mining laws, the USFS states that effects to the IRA

and its potential wilderness values are protected because no new roads are to be built and

minimal overland travel will occur. In a discussion with the USFS Geologist in their Tonopah

office he could only say that any decision to approve a mining plan would be made in

Washington. The project site has good existing road access but currently any approval of new

roads within the IRA would be a significant issue.

The entire subject of Roadless Areas has given rise to a great deal of litigation over the past

several years, which may have ended in October 2012 with the Supreme Court refusing to hear

an appeal brought by the state of Wyoming and the Colorado Mining Association which sought

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to have the entire Roadless Area designation overturned because the Forest Service had

created de facto wilderness areas, which was argued could only be enacted by Congress.

We have researched the issue of mines approved in IRAs in recent years, and so far have found

none that have been approved or denied, although a number of exploration projects have been

approved including some in Nevada (almost always referred to in news articles as “mining

projects”). It is important that the Company deal with this issue in a proactive manner and

begin the process to build support for the project locally and at the state level. Developments

regarding this issue should be monitored as it has the potential of becoming an obstacle to USFS

approval of a plan of operations.

11.4 Greater Sage-grouse

Another issue that potentially could have a significant impact on the permitting schedule and

costs, and in the worst case might prevent USFS and/or BLM approval of the Plan of Operations

is the potential listing of the Greater Sage-grouse as an endangered species.

The greater sage-grouse has been for several years subject to efforts by environmental groups

to see that this it is placed on the Endangered Species List. At this time the environmental

groups appear to be nearing success. In March, 2013 the US Fish and Wildlife Service

announced the availability of a final report dated February, 2013 entitled Greater Sage-grouse,

Conservation Objectives: Final Report. In this announcement the following statement is made:

“After a thorough analysis of the best available scientific information, the Fish and

Wildlife Service has concluded that the Greater Sage-grouse warrants protection under

the Endangered Species Act. However, the Service has determined that proposing the

species for protection is precluded by the need to take action on other species facing

more immediate and severe extinction threats.”

This statement recites the same finding made in the Federal Register March 23, 2010. (This

Federal Registry entry discusses the sage-grouse situation extensively in 105 pages.) The Fish

and Wildlife Service is not moving expeditiously on this finding but may formally propose the

listing in 2014 or 2015.

The US BLM issued a document in November 2013 entitled Nevada and Northeastern California

Greater Sage-Grouse Draft Land Use Plan Amendments and Draft Environmental Impact

Statement. (Notice of availability published in the November 1, 2013 Federal Register). If this

Land Use Plan were to be adopted as proposed, severe restrictions would be placed on mineral

development in Nevada and many projects now contemplated would be unable to obtain

needed governmental approvals. The American Exploration and Mining Association (formerly

the Northwest Mining Association) is preparing extensive comments on numerous technical and

legal aspects of this document. If sufficient changes are not made by BLM, litigation is expected.

This issue should be closely monitored by Star Gold.

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The Nevada Department of Wildlife (NDOW) has produced a map entitled Greater-sage grouse

Habitat Categorization which as nearly as can be determined on a map of this scale

(approximately 1 inch = 39 miles on the copy below) places the Longstreet Project near habitat

areas classified as Habitat of Moderate Importance or Low Value Habitat/Transitional Range.

Depending on the exact project location it may be just outside the Population Management

Unit (PMU) boundary. A downloaded copy of this map is attached (a much better copy is

available on the internet as noted in the key on the page following the map.)

The site specific baseline studies of the Longstreet site will better define the actual population

of sage-grouse (if any) and suitability of habitat, it is important that early contact be made with

NDOW to see if any more information has been developed for the Longstreet Project region and

to be sure that the baseline study design meets with their approval. Early discussions of the

sage-grouse issue should also be held with the BLM and USFS, again to review study design.

At this stage of the development the Greater Sage-grouse issue appears to not have a

significant impact on the project. No other endangered, or proposed-to-be-endangered,

species are known to exist on the Longstreet site.

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Figure 11.1 Greater Sage-grouse habitat

Longstreet

Project

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No warranty is made by the Nevada Department of Wildlife as to the accuracy, reliability, or

completeness of these data for individual use or aggregate use with other data.

This map is available for download at www.ndow.org/wild/conservation/sg

11.5 Cultural resources

The site of the Star Gold claims has seen previous mining operations dating back to

approximately 1904 and concluding in 1929. Small scale milling operations were conducted on

the site and a small community is known to have existed in close proximity to the old mine.

Little remains of the past operations or community but there are the ruins of a cabin thought to

have belonged to the original prospector, Mr. Longstreet, within the project boundaries. There

are also some non-significant Native American camp locations and artefacts within the project

boundaries. If the cultural resources assessment designates any of these areas as worthy of

preservation the proposed mining and leaching activities will not be impacted.

11.6 Environmental and permitting conclusions

The permitting of mining operations within the United States is never a simple process and is

always time consuming and expensive. Nevada is considered one of the most favourable

jurisdictions to permit a heap leach gold operation due to the long history of operations in the

state. Given the current information regarding this project it is estimated that the permitting

timeline for this project could be two to four years. This is largely due to having to permit

operations on USFS and BLM lands, which may be mitigated somewhat by the modest size of

the proposed operation.

The permitting timeline can be reduced by proactively proceeding with the likely required

baseline studies that include hydrology, flora and fauna and cultural resources. These studies

could largely be accomplished during the continued advanced exploration and engineering

phase of the project. The USFS and BLM as well as local experts should be consulted on specific

studies and their scope before work is undertaken so as to optimize this effort. The IRA issue

needs to also be proactively dealt with to get clarity on what developments will or will not be

allowed to occur on the site.

AMPL concludes that there are no recognized potential environmental or permitting fatal flaws

regarding this project.

______________________________________________________________________________________________


12 PROJECT DEVELOPMENT SCHEDULE

The schedule for developing a mine at Longstreet remains uncertain. Figure 13.1 provides a

timeline for additional engineering studies, the EIS and permit acquisition, project construction

and commissioning to reach commercial production in three years. Opportunities exist to fast-

track the Project.

This Scoping Studywould be followed by a Preliminary Feasibility Study which would necessitate

additional data collection, broader field investigations and more detailed engineering, to

address the major issues identified in this study, while ensuring study expenditures are

optimized. A PFS for a project of the scope of Longstreet can be expected to require 6-12

months, depending on the amount of data that is required to be collected.

Following the delivery of a positive PFS, time and funding must be sought to complete a

Feasibility Study to the standards demanded by mine financiers. The Feasibility Study could take

from six months to a year to complete.

Processing equipment lead times will be on the critical path of constructing the ADR plant thus

consideration should be given to ordering long-lead time items as early as possible.

Investigation of a modular ADR plant to suit the processing throughput criteria of the

Longstreet Project is recommended

The construction period for the Longstreet Project will be relatively short. Main construction

components would be earthworks (site road construction, leach pad foundation, pond dams,

ROM pad). Additional construction activities will include the mining equipment maintenance

facility (by the mining contractor), office structures, services, installation of the leach pad liner

and the ADR plant.

______________________________________________________________________________________________Star Gold Inc. Longstreet Project Scoping

Study 76

Figure 12.1 Longstreet Gold Project engineering and development schedule

Year Year 1

Activity Month 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37Initiate PFSPFS study

In-fill core drillingSample & assay programRevise resource estimate

Geotechnical drillingGeotechnical testing

Metallurgical testworkProcess engineering

TSF designMining study

EIAEIS filed with Regulators

PFS reportPFS completeFinance DFSFS studyFS deliveredExecution financingProject constructionCommisioning periodCommercial production

Year -1Year -2Year -3

______________________________________________________________________________________________


13 CAPITAL EXPENDITURES ESTIMATES

13.1 Basis for estimates

The capital expenditures estimates are based on budget pricing from suppliers for critical components, consultants, contractors and a review of other projects. Smaller equipment and

facilities component costs were factored based on industry norms for the type of facility being

constructed and, where possible, adjusted to reflect local conditions.

Capital expenditure estimates have an accuracy of +- 50%.

All expenditure estimates are in 2014 constant US Dollars.

13.2 Mining

Mine capital expenditures are primarily related to mine services. The total mine pre-production expenditures are expected to be approximately $US0.22 million. These expenditures are

included in the mine and surface services infrastructurecosts as they are mainly related to site

roads and power. No pre-stripping expenditures are included as the initial mineralized material can be accessed directly. All mining equipment and related facilities would be provided by a

contractor.

No mine sustaining capital expenditures are envisaged because of the short mine life.

A contingency of 15 percent is included in the capital expenditures estimate.

13.3 Heap leach and processing plant

The processing capital cost estimate covers the design and construction of the heap leach and

ADR plant, together with certain on-site and off-site infrastructure.

A contingency of 15% was incorporated into the total cost of the project for the pre-production

expenditures.

The initial heap leach pad construction capital expenditures total $US2.25 million. The

expenditures would be for the heap pad and liner, initial drip piping, pregnant solution

collection and water diversion and collection ditches around the base of the pad. The

construction of a gold recovery plant would require capital expenditures of approximately

$8.10 million. Table 13.1 presents the ADR plant capital expenditures.

For the processing plant, equipment pricing is based on the equipment list, which is generated

from the process flow diagram.

Other direct costs (eg. earthworks, concrete, structural, piping, electrical, instrumentation, etc.)

were factored based on the total installed cost of process equipment

______________________________________________________________________________________________


Table 13.1 ADR Plant Capital Expenditure Estimate

13.4 Infrastructure and support facilities

The costs for the infrastructure are primarily allowances based on in-house experience from

other similar projects. The cost in the estimate is based on the assumption of a 2-year starter

pad. This area needs input from a specialist geotechnical company at the next stage of the

project to develop more accurate costs.

Total pre-production capital expenditures for project infrastructure and surface department are

estimated to be approximately $US4.2 million. Table 13.2 provides the infrastructure and

support services capital expenditures breakdown. Major expenditure components are for

water supply, power generation and an office/shop/warehouse complex.

Area TOTAL

($US)

Lime Bin $175,000

Carbon Absorption $1,189,364

Elution & Carbon Regeneration $2,183,381

Electrowinning & Refinery $357,388

Reagents $185,000

Services, Air & Water, Tanks, RO Plant $314,620

Processing Plant Equipment Sub-total $4,404,753

Factored Direct Costs Sub total $3,692,124

Total ADR Expenditures $8,096,877

______________________________________________________________________________________________


Table 13.2 Infrastructure Capital

13.5 Owners costs

Project in-directs are all estimated expenditures for the project (such as EPCM and project

management personnel) borne directly by Star Gold in completion of the project construction.

These estimated expenditures total $US4.54 million (including US$2.00 million for permitting)

over the 1 year pre-production period. Owner`s costs also include all equivalent General and

Administration costs which would be incurred during the construction phase.

13.6 Total capital expenditures

The estimated project pre-production capital expenditure, inclusive of contingencies and

working capital, is approximately $US25.4 million. The total expenditures include EPCM,

contractor overheads and a 15% contingency on all estimated expenditures. A summary of

project pre-production capital expenditures is presented in Table 13.3. A working capital allowance of $US3.7 million, representing 3 months operating costs, is estimated to be required.

Component Total Cost

($US million)

Site Preparation 0.097

Access Roads 0.162

Process Water 2.000

Water Reclaim 0.202

Power Supply 0.900

Fuel Storage & Distribution 0.081

Water & Sewage Treatment 0.081

Service Complex Buildings 0.324

Water Supply & Distribution 0.162

Mobile Equipment/Power Supply 0.162

Communication 0.049

Total Infrastructure Expenditures $4.220

______________________________________________________________________________________________


Table 13.3 Project Pre-Production Capital Expenditures

Cost Component Expenditure

($US)

Permitting $ 2,000,000

Mine $ 220,000

Heap Leach Pad Processing Plant

$ 2,250,000 $ 8,097,000

Surface Infrastructure & Mobile Equipment $ 4,000,000

EPCM, Contractor O/H & Owners Costs Contingency

$ 2,535,000 $ 2,565,000

Total Capital Expenditures $21,667,000

Working Capital $ 3,690,000

TOTAL EXPENDITURES $25,357,000

The capital estimates include the following conditions and exclusions:

• The crushing plant and supporting infrastructure capital expenditures are not

included in the capital cost estimate as it will be provided by the mining contractor;

• Qualified and experienced construction labour will be available at the time of

execution of the project;

• There is no detailed geotechnical and drainage assessment of the site, therefore no

allowance for special ground preparation has been made;

• A water supply capable of supplying the required demand of the processing plant is

assumed to be available;

• No extremes in weather have been anticipated during the construction phase; and

• No allowances have been included for construction-labor stand-down costs.

13.7 Sustaining capital

No sustaining capital expenditures are estimated because of the relatively short mine life.

13.8 Closure costs

Closure costs have been estimate at $1 million at the end of the project life, shown on the cash

flow model as a reduction in working capital credit.

______________________________________________________________________________________________


14 OPERATING COST ESTIMATES

14.1 Basis for estimates

Operating costs are based on US and other country norm prices from suppliers and other similar type projects, for consumables and parts. The cost of power is based on diesel generated power. Critical operating cost components are based on the following costs:

The diesel fuel price is assumed to be $0.94 / litre.

The electrical power cost is assumed to be $0.22 per kWh. Labour costs for the operating period are based on the manpower schedules presented for each

department and the associated labour costs. The costs include a burden component of

approximately 35 percent. Labour rates are based on local rates where available and/or

contractor costs in the region and country, for similar types of work. Where costs were either

not available or irrelevant, costs from other similar projects were used. The rates used include

all cost and profit components payable to contractors.

All costs are quoted in constant 2014 US Dollars.

14.2 Mining

The mine operating cost estimates were developed from a cost base of similar types of projects and conditions. The average total mine operating costs are estimated to be $US9.09 per ton of potentially economic mineralization. Potentially economic mineralization unit mining costs are estimated to be $US7.00 per ton, which includes trucking to and dumping at the heap leach pad and waste unit mining costs are approximately $3.00 per ton.

14.3 Heap leach and gold recovery plant

The heap leach operating cost includes crushing, stacking the leach pad, installation and repair of drip piping, reagents for leaching and collection and pumping of pregnant solution to the gold recovery plant. The gold recovery plant costs comprise gold absorption from pregnant solution systems, gold electro-winning and refining costs, carbon regeneration and return of cyanide solution to the heap leach operation. The total operating cost would be approximately $US3.65 per ton of potentially economic mineralization. A breakdown of the cost is presented in Table 14.1 and includes labour, consumable supplies, electrical power usage, maintenance supplies, and other applicable costs.

______________________________________________________________________________________________


Table 14.1 Processing Plant Operating Cost

Function Unit OPEX

($US/t)

Labour – Metallurgy & Production 0.81

Labour - Maintenance 0.28

Power 0.68

Maintenance Materials 0.24

Reagents & Consumables 1.33

Miscellaneous 0.31

TOTAL $3.65

The operating costs for the processing plant are based on the following criteria:

Labour: Around the clock operations are based on a two shift rotation of 12-hour

shifts using a 7 days on, 3 days off, 7 days on 4 days off cycle. Non-shift labour is

based on a 40-hour work week, working 5-8 hour shifts.

The man power costs for this Project were estimated using other mining projects

in Western United States, based on labour rates and payroll burdens (35% overhead).

Commodity usage rates were developed from recent test work. Unit pricing for

commodities was taken from a data base of similar projects. No premium or transportation fees on the costs of consumables were added. This will need to be

investigated and discussed with reagent vendors during the next phase of the project.

Electrical power consumption and estimates were based on equipment connected

loads. A factor of 80% was used to estimate the operating load from the connected load.

Maintenance supplies for stationary equipment are based on 3.5% of installed

mechanical and electrical equipment costs. For piping, electrical, and instrumentation, a factor of 1.5% was used to estimate maintenance supplies.

Factor rates are based on experience.

The crushing plant will be operated by contractors who will be responsible for

providing electrical power, staffing and consumables required for operating the

plant. As such the estimated operating costs do not incorporate any costs associated with the crushing plant.

______________________________________________________________________________________________


14.4 General & Administration Costs

The estimates for G&A costs encompass all operating costs associated with operating the offices and providing materials and supplies for staff functions. Administration operating costs

include costs and taxes for maintaining the property in good standing, land taxes, and resource usage fees (water, etc.).

The total yearly G&A costs are estimated to be approximately $2.0 million (presented in Table 14.2), of which approximately $1.0 million is for salaries and benefits. Employee burdens

account for approximately 35% of the total salary for each employee.

Annualized site G&A costs are estimated at $2.01 per ton of ore processed. However, the life-

of-mine G&A cost will be $2.13 per ton as a result of the partial final year of operations and

fixed costs to maintain production.

______________________________________________________________________________________________


Table 14.2 General and Administrative Operating Cost Components

Component Annual

Cost

($US)

Salaries & Overhead $951,000

Training $10,000

Safety Equipment $5,000

Medical, Health & Safety $50,000

Government Relations $20,000

Power $296,000

Travel & Accommodations $20,000

Marketing $25,000

Legal and Accounting $30,000

Consultants $150,000

Shipping, Courier and light freight $30,000

Communications $25,000

Office Supplies $15,000

Computer Supplies $20,000

Light Vehicles Operation $25,000

Roads and Yards Maintenance $30,000

Insurance $100,000

Human Resources $30,000

Bank Costs $10,000

Surface ITC $50,000

Buildings Maintenance $5,000

Electrical Distribution Repair $5,000

Water Supply & Water Treatment $50,000

Office Equipment Leases $12,000

Security Supplies $5,000

Cleaning contract $20,000

Dues & Subscriptions $5,000

PR $20,000

TOTAL G&A COSTS $2,014,000

______________________________________________________________________________________________


The mine management and administration roster and costs have been estimated in Table 14.3. A total of 13 people will be employed in this area, most of which will be staff positions. They will be responsible for the management, administration, personnel, accounting, purchasing needs, and distribution of material to the operation, site security, health and safety, and environmental issues. The total costs for G&A labour is US$0.95 per ton of ore processed.

Table 14.3 G&A Manpower Costs

14.5 Dore transport and refining charges

Transport and refining costs of $US4.00 per ounce gold have been included in the cashflow model and are based on relative norms.

14.6 Project total operating costs

The estimated total average operating cost (excluding smelting and refining) for the mine is approximately $14.87 per ton of potentially economic mineralization. Table 14.4 presents a summary table of life of mine average operating costs for each department on a cost per ton of potentially economic mineralization.

Position Complement Annual Fringe Total

Salary Benefits Cost

($US) 35% ($US)

Mine Manager 1 125,000 35% $169,000

Senior Engineer 1 66,600 35% $90,000

Accountant 1 52,000 35% $70,000

Eng/Geo technicians 2 55,000 35% $149,000

Purchasing/Warehouse Manager 1 70,000 35% $95,000

Environmental Coordinator 1 62,400 35% $84,000

Medical Contract 1 52,000 35% $70,000

Security Guard 4 31,200 35% $168,000

Site Services 1 41,600 35% $56,000

Grand Total 13 $951,000

______________________________________________________________________________________________


Table 14.4 Project Operating Cost Summary

Department Cost

($US/t Mined)

Mine $ 9.09

Processing & Environmental $ 3.65

Surface Dept. and G&A $ 2.13

TOTAL $14.87

14.7 Exclusions

For the purpose of this study, value added taxes and other taxes, along with import duty costs, have not been included. Crushing costs along with transportation and refining charges for gold

bullion bars are not included in the operating costs but are considered in the financial model as

are rehabilitation costs (included in deferred capital schedule), land tenure and claim fees,

exploration costs and all costs associated with areas beyond the property limits.

______________________________________________________________________________________________


15 ECONOMIC ANALYSIS

15.1 Basis for analysis

The expected base case cash flow estimates have been made using a forecast long-term gold price of $US1,350 per ounce. A summary of the expected parameters used for the financial analysis is presented in Table 15.1.

Table 15.1 Cash Flow Model Inputs

Component Parameter

Potentially Mineable Resource (Indicated & Inferred), including mining dilution & recovery

4.4 million tons

Estimated Mining Dilution 5 percent @ 0% grade

Average mill head grade, gold 0.022 opt

Average mill head grade, silver 0.53 opt

Payable gold 82,450 ounces

Payable silver 348,200 ounces

Average long-term gold price $1,350 per ounce

Average long term silver price $24.00 per ounce

Pre-Production Capital including Working Capital $25.4 million

Total Sustaining Capital $0

Royalty 3% NSR

Closure Cost $1 million

Estimated Operating Costs ($/ton) $14.87

Life of Mine 4.4 Years

The cash flow analysis has been conducted on the assumption of 100% equity investment and excludes any element or impact of financing arrangements. All exploration and acquisition

costs incurred prior to the production decision are also excluded from the cash flows.

Capital expenditures, as shown in the capital section, would be incurred over a one year period,

which is reflected in the discounted cash flow calculations. The cash flows include sustaining capital and capital expenditures contingency of approximately 15%.

Revenue is based on payments for gold by gold refiners. Costs for metal sales and shipping are

included in the deductions that the refiner makes.

The expected cash flow analysis used the metal prices indicated above. The discounted cashflow analysis has been based on 2014 Constant US Dollar values.

______________________________________________________________________________________________


15.2 Metal price derivation It is common practice to consider the long-term average price of gold when deriving a price to evaluate a

mineral deposit. Neither AMPL nor Star Gold are able to forecast the price of gold.

The price of gold has exhibited strong variability for some years, rising until mid-2011, fluctuating above

$1,600 per ounce until the end of 2012, then eroding to its current level in the $1,300 - $1,400 per ounce

range. The trends and 36 month moving average gold price (red line in graph) shown in Figure 15.1

present the gold price variability since 2006 and underscore the selection of a US$1,350 price per ounce

of gold for this evaluation of the Longstreet Project.

Figure 15.1 Gold price trend 2006 -2014

Courtesy Kitco

15.3 Financial returns

The levels of accuracy for this study are +/- 30% to 50%. This Scoping Study relies on Indicated

Mineral Resources but also Inferred Mineral Resources. Inferred Mineral Resources are

considered too speculative geologically to have economic considerations applied to them that would enable them to be categorized as Mineral Reserves.

______________________________________________________________________________________________


The summary cashflow model for the Longstreet project is presented in Table 15.2 using the expected project parameters.

The expected investment and returns based on the estimated cash flow for the Project are shown in Table 15.3.

______________________________________________________________________________________________


Table 15.2 Longstreet Gold Project Cash Flow Model

Description Unit Unit Rate Total

-1 1 2 3 4 5

Mineable Resources

Start of Period tons 4,418,183 4,418,183 3,418,183 2,418,183 1,418,183 418,183

Mined tons 0 1,000,000 1,000,000 1,000,000 1,000,000 418,183 4,418,183

End of Period tons 4,418,183 3,418,183 2,418,183 1,418,183 418,183 0

Production

Mineralization Mined tons 1,000,000 1,000,000 1,000,000 1,000,000 418,183 4,418,183

Stripping Ratio 0.7 0.7 0.7 0.7 0.7

Waste Mined tons 697,351 697,351 697,351 697,351 291,620 3,081,023

Mineralization to Leach Pad tons 0 1,000,000 1,000,000 1,000,000 1,000,000 418,183 4,418,183

Grade gold gpt Au 0.022 0.022 0.022 0.022 0.022 0.02

Grade silver gpt Ag 0.53 0.53 0.53 0.53 0.53 0.53

Heap Leach/ADR Gold Recovery % 86% 86% 86% 86% 86% 86% 86%

Heap Leach/ADR Silver Recovery % 15% 15% 15% 15% 15% 15% 15%

Gold Produced Ounces 0 13,997 18,662 18,662 18,662 12,470 82,452

Silver Produced Ounces 0 59,108 78,810 78,810 78,810 52,659 348,197

Revenue

Gold price $/oz $1,350 $1,350 $1,350 $1,350 $1,350 $1,350 $1,350

Silver price $/oz 24.00 $24 $24 $24 $24 $24 $24

Gold Revenue $ $18,895,000 $25,194,000 $25,194,000 $25,194,000 $16,834,000 $111,311,000

Silver Revenue $ $1,418,580 $1,891,440 $1,891,440 $1,891,440 $1,263,827 $8,356,727

Transport & Refining $/oz $4.00 $0 $56,000 $75,000 $75,000 $75,000 $49,878 $330,878

Net Revenue $ $0 $20,257,580 $27,010,440 $27,010,440 $27,010,440 $18,047,949 $119,336,849

Operating Costs

Mine - O/P Ore $/t $7.00 $0 $7,000,000 $7,000,000 $7,000,000 $7,000,000 $2,927,000 $30,927,000

Mine - O/P Waste $/t $3.00 $2,092,000 $2,092,000 $2,092,000 $2,092,000 $875,000 $9,243,000

Heap Leaching & Gold Recovery $/t $3.65 $0 $3,653,000 $3,653,000 $3,653,000 $3,653,000 $2,440,871 $17,052,871

General & Administration $ $2,014,000 $2,014,000 $2,014,000 $2,014,000 $2,014,000 $1,345,720 $9,401,720

Total Operating Cost $ $0 $14,759,000 $14,759,000 $14,759,000 $14,759,000 $7,588,591 $66,624,591

Operating Profit $0 $5,498,580 $12,251,440 $12,251,440 $12,251,440 $10,459,358 $52,712,258

Royalty - MinQuest 5 3% $0 $164,957 $367,543 $367,543 $367,543 $313,781 $1,581,368

EBITDA $0 $5,333,623 $11,883,897 $11,883,897 $11,883,897 $10,145,577 $51,130,890

Capital Expenditures

Permitting $ $2,000,000 $2,000,000 $2,000,000

Mine & Surface Services Infrastructure $ $2,219,791 $0 $2,219,791

Process Water $ $2,000,000 $2,000,000 $2,000,000

Indirects & Project Management $ $2,535,093 $0 $2,535,093

Heap Pad Construction $ $2,249,562 $2,249,562

Gold Recovery Plant $ $8,096,877 $8,096,877

Contingency $2,565,198 $2,565,198

Working Capital $ $3,689,750 -$3,689,750 $0

Mine Closure $ $1,000,000 $1,000,000

Total Capital Expenditures $ $21,666,522 $3,689,750 $0 $0 $0 -$2,689,750 $22,666,522

Project Pre-Tax Cashflow $ -$21,666,522 $1,643,873 $11,883,897 $11,883,897 $11,883,897 $12,835,327 $28,464,368

Project Cumulative Cashflow $ -$21,666,522 -$20,022,649 -$8,138,752 $3,745,144 $15,629,041 $28,464,368

IRR 29%

NPV 5% $19,788,000

10% $13,331,000

15% $8,469,000

Year

______________________________________________________________________________________________


Table 15.3 Cash Flow Summary

Undiscounted Net Revenue $119 million

Undiscounted Cashflow $ 29 million



IRR 29%

Payback Period 2.7 years

15.4 Sensitivity analysis

Sensitivity analysis was performed for metal prices, capital expenditures, operating costs, mined grades and heap leach recoveries using 15 percent positive and negative variations, except for recoveries which were limited to +/- 5% (as this is realistically the range that recoveries could be expected to vary). The project is very sensitive to changes in metals prices and reasonably sensitive to changes in all the other variables. The results of the sensitivity analysis are presented in Table 15.4.

Table 15.4 Pre-Tax Sensitivity Analysis

The IRR and NPV sensitivities to variations in key parameters are depicted graphically in Figures

15.1 and 15.2. The IRR is most sensitive to variations in metal prices and mined grades and least

sensitive to operating costs. Potential expected metals recoveries variations show limited

Variable Variation NPV @

10% ($millions)

IRR (%)

Metal Prices +15% 25 46

Metal Prices -15% 1 12

Capital Expenditure +15% 10 24

Capital Expenditure -15% 16 36

Operating Costs +15% 6 19

Operating Costs -15% 20 40

Mined Grades +15% 25 45

Mined Grades -15% 1 12

Metals Recovery +5% 17 35

Metals Recovery -5% 9 24

______________________________________________________________________________________________


sensitivity but should the recoveries fall to a greater percentage the viability of the operation

could quickly be rendered uneconomic.

Figure 15.2 Pre-Tax IRR Sensitivities

Figure 15.3 NPV10 Sensitivity Analysis

An isolation of the effect of gold price variability on the Longstreet Project provides a clear picture of the

impact of this variable, as shown in Figures 15.4 and 15.5.

0

10

20

30

40

50

60

-15% -10% -5% 0% 5% 10% 15%

IRR

(%

)




-5

0

5

10

15

20

25

30

35

-15% -10% -5% 0% 5% 10% 15%

NP

V (

$U

S m

illio

ns)




______________________________________________________________________________________________


Figure 15.4 Project IRR Sensitivity to Gold Price

Figure 15.5 Project NPV Sensitivity to Gold Price

______________________________________________________________________________________________


15.5 Economic interpretations and conclusions

Based on the study results, conclusions are:

1. The Project provides positive and robust returns. 2. Longstreet is a small deposit which can be developed for production at reasonable

cost in a near-term horizon, providing regulatory permits are achieved.

3. The Project is most sensitive to variations in the price of gold and variations in the mined grade of mineralized material.

4. Increasing the tonnage delivered to the heap leach pad by discovering and mining

economic satellite deposits also has a significant positive impact on project returns.

The initial capital investment will be repaid by the Main Zone and almost all of the

operating profits from other deposits would report to the cashflow line.

______________________________________________________________________________________________


16 PROJECT RISK ASSESSMENT

The Longstreet Project is technically uncomplicated because of the near surface nature of the deposit and relatively simple open pit mining. The heap leach system is well proven for these

types of gold mineralization in Nevada and should achieve good gold recoveries. Infrastructure

requirements are also relatively risk free as the mine is in an area of other economic activity with many regional services.

The main risks to project success would be:

Gold price variations, particularly if gold price drops by more than 15% from the

$1,350 per ounce level;

Water supply is a major component which requires further work to identify

sources and adequate volume;

The confidence in the mineral resource represents a risk to the project. Steps

should be taken to upgrade Inferred Resources to at least Indicated Resource

category;

Environmental risks, particularly the ability of the project proponent to secure

permits for road building in an area deemed to be a Roadless Area;

Pre-production capital expenditures represent a relatively low risk as the mine

development and surface infrastructure required to commence production are not overly extensive. Regional communities provide much of the support services for

employees and the mine.

______________________________________________________________________________________________


17 CONCLUSIONS AND RECOMMENDATIONS

17.1 Conclusions

This Scoping Study has identified a potentially mineable resources of 4.4 million short tons at

0.022 ounces Au per short ton and 0.53 ounces Ag per short ton. The deposit would be mined

by the open pit mining method with gold and silver extracted by heap leach and a gold recovery

plant. The mine site infrastructure facilities would be minimized but include a small surface

shop, warehouse, office complex and water treatment facility.

The mine would operate at 1 million tons per annum and produce approximately 18,700 ounces

of gold and 78,800 ounces silver. Gold and silver recoveries would be 86 percent and 15

percent respectively.

Economic analysis has indicated a robust return depicted by a positive NPV and an Internal Rate

of Return estimated at 29%.

The IRR is most sensitive to variations in metal prices and mined grades and least sensitive to

operating costs. Potential expected metals recoveries variations show limited sensitivity but

should the recoveries fall to a greater percentage the viability of the operation could quickly be

rendered uneconomic.

Based on the study results, the conclusions of AMPL are:

1. The project provides positive and robust returns.

2. Longstreet is a small deposit which can be developed for production at reasonable

cost in a near-term horizon, provided regulatory approval and permits are acquired.

3. The mined grade of ore is an important variable for the success of the operation as is

mining cost. Operating management efforts during mine production must be

focussed on these two parameters.

4. The Project is most sensitive to variations in the price of gold and variations in the

mined grade of mineralized material.

5. The economics of the project would be improved with the discovery and exploitation

of economically viable satellite deposits. Once the capital investment has been

repaid by the Main Zone the operating profits from other deposits would enhance

the Project cash flow.

6. Water sourcing is the largest technical risk factor, particularly to capital expenditures

and operating cost estimates. Ideally a well source will be identified and thus avoid

the added cost of piping water to the site from Five Mile Spring on Clifford Ranch,

currently the nearest identified water source, 12 miles from the Project site. It is not

known if Five Mile Spring produces adequate volume nor if the owner of Clifford

Ranch would agree the sale of water from the spring. Preliminary investigations have

identified two potential sources of well water; one mile to the NE in the Monitor

Range and approximately five miles east in Stone Cabin Valley.

______________________________________________________________________________________________


7. AMPL has reviewed the permitting requirements of the US Forest Service, The

Bureau of Land Management and the Nevada Division of Environmental Protection

and estimates that, without objection during the public disclosure period of

permitting, the Longstreet Project will require from two to four years to secure the

permits required to begin constructing and operating the mine.

17.2 Recommendations

Based on the results of this Scoping Study, recommendations follow:

Geology

Recommendations for the next phase of mineral resource estimation include:

1. Consider a drilling program to explore mineralization at depth and test the nearby

Central Ridge mineral occurrence.

2. Consider further drilling to better understand the transition zone between oxide &

sulfide to determine the maximum extent of leachable gold mineralized material.

Mining

1. Undertake geotechnical work for open pit slope angles optimization.

2. Obtain firm quotations from qualified local mining contractors

Heap Leaching & Processing Plant

1. Conduct column test work on the oxide adit material to test the mineralogical variability of the deposit.

2. Conduct column leach tests using finer material in conjunction with high pressure rolls i.e. P80 ¼-inch (6.3mm) in order to maximize silver recovery.

3. Use 60 days column leach time for the next phase of testwork as the leach kinetics for gold are fairly rapid and the silver recovery did not increase dramatically even after 190 days of leaching.

4. Conduct column tests using site water as opposed to laboratory tap water in order to determine the effects of site water on leach kinetics.

5. A HPGR (high pressure grinding rolls) evaluation should be considered to investigate improved silver recovery on the master blend composite ore (generic tests show that HPGR use often leads to the formation of micro cracks in the ore which may improve silver leaching kinetics). This would require a Static Pressure Test (SPT) to be performed.

6. Load/permeability tests are recommended on column leach residue samples to confirm permeability under compressive loading.

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Infrastructure

1. A hydrological study is recommended to identify proximal water sources of adequate

volume to sustain the Longstreet operation.

Environment and Permitting

1. Initiate baseline studies as soon as possible as a precursor for applications for

permits to construct and operate the Project.

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18 REFERENCES

1) Agnerian, H, Routledge, R. E., and Gharapetian, R, December 2013, Technical Report On

The Longstreet Gold-Silver Property, Nevada, Prepared For Star Gold Corporation by Agnerian

Consulting LTD, North York, Ontario, Canada

2) Eakin, T. E, 1962, Groundwater Appraisal of Ralston and Stone Cabin Valleys, Nye

County, Nevada, Groundwater Resources-Reconnaissance Series-Report 12, U.S. Department of

the Interior, U.S. Geological Survey and State of Nevada Department of Conservation and

Natural Resources

3) Heilweil, V.M. and Brooks, L. E, editors, 2010, Conceptual Model of the Great Basin

Carbonate and Alluvial Aquifer System, Scientific Investigations Report 2010–5193, U.S.

Department of the Interior, U.S. Geological Survey.

4) Price, J. G., 2003, Geology of Nevada, Preprint from Castor, S.B., Papke, K.G., and

Meeuwig, R.O., eds., 2004, Betting on Industrial Minerals, Proceedings of the 39th Forum on the

Geology of Industrial Minerals, May 19–21, 2003, Sparks, Nevada: Nevada Bureau of Mines and

Geology Special Publication 33.

5) Rush, F.E. and Everett, D.E. 1964, Groundwater Appraisal of Momitor, Antelope and

Kobeth Valleys, Nye County, Nevada, Groundwater Resources-Reconnaissance Series-Report 30,

U.S. Department of the Interior, U.S. Geological Survey and State of Nevada Department of

Conservation and Natural Resources

6) Smyth, R. C. and Sharp, J.M., 2006, The hydrology of Tuffs, Special Paper 408, Geologic

Society of America.

7) Noland, P., February 16, 2014, Longstreet Project, Nye County, Nevada Revised

Technical Review and Resource Estimate

8) Kantor, J.A., 2014, Longstreet Gold Project Exploration Potential, Star Gold Study, 2014

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APPENDIX I

2013 Metallurgical Testwork Report

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APPENDIX II

SITE WATER BALANCE

______________________________________________________________________________________________Star Gold Inc. Longstreet Project Scoping

Study 103

Longstreet Project Water Balance evaporation

precipitation assumed 5 % = 21 m3/h

make up water 184 m3/h

required 21 m3/h 184 m3/h

evaporation

Mine fresh water consumption

? m3/h LEACH PAD

Mine Pit

Surface Pit

Water Water

evaporation

Solution loses 10.5 m3/h

Pit dewatering Environment ? precipitation

cell Solution loses 10.5 m3/h

evaporation Makeup water

21 m3/h

seepage ? precipitation evaporation

368 m3/h

PLS Pond BLS Pond

Event Pond

173.5 m3/h

ADR Plant173.5 m3/h

Total Fresh Water Demand

make up water - 21 m3/h

booster pump Potable water - 1 m3/h

Reagents - 1 m3/h

Dust supression - 1 m3/h

Well # Well # Well #

BLS Tank

Process water Tank

Reverse Osmosis

SCOPING STUDY - Star Gold€¦ · This Scoping Study has identified potentially mineable resources...

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