reprocessing the bangalee creek tailings - jason downes - 2012

UNIVERSITY OF GUYANA

FACULTY OF TECHNOLOGY

DEPARTEMENT OF GEOLOGICAL ENGINEERING

Reprocessing the Bangalee

Creek Tailings:

A Preliminary Feasibility

Study

Final Research Paper

Submitted By

Jason Downes

October, 2012



Reprocessing the Bangalee Creek Tailings: A Preliminary Feasibility Study

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This project is submitted in partial fulfilment of the requirements for the Bachelor of

Engineering

Geological Engineering

A Research Submitted by:

JASON DOWNES

10/0939/1647

DEPARTMENT OF GEOLOGICAL ENGINEERING,

FACULTY OF TECHNOLOGY,


Internal Supervisor:

Mr Sherwood Lowe

External Supervisor:

Mr Andre Archer

Manager

Metallica Commodities Corp Guyana Inc.

October, 2012




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ABSTRACT

Encouraged by the escalation of gold price experienced within the past few years, the area of

tailings reprocessing and treatment has emerged as a brilliant, seemingly new avenue for

companies seeking to engage in mining without the high capital investment. There exist many

worldwide examples of companies (large, medium and small) actively engaged in this practice.

Locally Metallica CC had decided to enter into this avenue of mining, with them seeking to

reprocess tailings material located within Bangalee Creek Area, Omai. The aim of this research

was to conduct a preliminary feasibility study tailings dumps within the area.

A total of 208 samples were collected from 10 tailings stockpile within the project area.

Sampling was done using series regular grids with samples being retrieved at grid line

intersections at 1 m depth interval to the maximum depth of the tailing dump. Analysis for gold

through fire assay with gravimetric finish was conducted on samples with samples grade

ranging from 0-7g/t for individual samples. Through the application of Rockworks 15 software

package for data management and analysis, 3D grade distribution model of individual tailings

were produced for visualization of results. Subsequent statistical analysis of assay results led to

the determination of average grade values for each tailing, which along which along with

material tonnage data made it possible to estimate the size of the potential gold resource.

A potential gold resource of 101.09 Oz was determined to be contained within 5810.54 tons of

tailings material. At $1600 USD/Oz, the resource is valued at $161,747.20 USD. Based on the

mining method proposed for the reprocessing project, an operational cost of $19,845 USD is

expected per month. Profitability of the project was determined to rely on two factor; the first

being how fast the tailing can be mined and secondly how much of the gold can be recovered.

As of such, the feasibility of the project would be determined by the performance of the

processing plant and the operation as a whole. The current configuration and performance of

the processing plant indicates the potential for a profitable operation, but only if high enough

recovery percentage levels are attained.




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ACKNOWLEDGEMENT

First and foremost, thanks to the Almighty God for his guidance, health and strength that

enabled the completion of this research.

The researcher would like to express gratitude to the management of Metallica Commodities

Corp. for giving him the opportunity to conduct this research through the provision of the

necessary resources needed for the completion of the program. Many thanks to manager

(Guyana), Mr. Andre Archer for his guidance, support and encouragement provided throughout

the period of this project. To staff member such as Mr. Orin Riddle, Derick Limerick and Lennox

Cummings, thanks are extended to you for the assistance and when necessary, leadership

offered during field work. Members of the sampling team, special thanks to you guys for the

great job down.

Special thanks are extended to Mr. Sherwood Lowe, Head of the Department of Geology, for his

insights and recommendations provided and for his general critical analysis of this project in

order to make it a success, again thank you. Words of appreciation are also extended to Mr.

Evan Persaud and Mr. Lennox Tucker, lecturers at the University of Guyana, for their guidance

during the conceptualization of this research

Thank you all.




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TABLE OF CONTENTS

Abstract ................................................................................................................................ ii

Acknowledgement ............................................................................................................... iii

List of Figures ........................................................................................................................ 1

list of Maps ........................................................................................................................... 2

List of Tables ......................................................................................................................... 2

1 Introduction ................................................................................................................... 3

1.1 Background ...................................................................................................................... 3

2 Objectives ...................................................................................................................... 6

3 Literature Review ........................................................................................................... 7

3.1 Previous Work .................................................................................................................. 7

3.1.1 Omai Gold Mines Limited – Final Report of West Omai ........................................... 7

3.1.2 Historic Mining Activities .......................................................................................... 8

3.2 Review of other related literature .................................................................................. 11

3.2.1 Sona Resources Corp – Blackdome Mine Tailings, British Colombia ...................... 11

3.2.2 Lake Victoria Mining Company, Singida, Tanzania ................................................. 11

4 Scope of Work ............................................................................................................. 12

4.1 Location and Access ....................................................................................................... 12

4.2 Geological Setting .......................................................................................................... 15

4.2.1 Regional Geology .................................................................................................... 15

4.2.2 Local Geology .......................................................................................................... 15

5 Methodology and Design Approach .............................................................................. 17

5.1 Data Acquisition ............................................................................................................. 17

5.1.1 Desk Study ............................................................................................................... 17

5.1.2 Site Characterization ............................................................................................... 17

5.1.3 Sampling Strategy and Technique .......................................................................... 18

5.1.4 Sample Preparation and Laboratory Analysis ......................................................... 20

5.2 Data Analysis .................................................................................................................. 21




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5.2.1 Processing and Analysis of Downhole Data ............................................................ 21

5.3 Evaluation ....................................................................................................................... 23

6 Results and Analysis ..................................................................................................... 24

6.1 Calculation of Tailing Average Au Grade ....................................................................... 24

6.2 Mineral Resource Estimation ......................................................................................... 33

6.3 Operational Cost Estimation .......................................................................................... 38

6.3.1 Mining Method ....................................................................................................... 38

6.3.2 Cost Assessment ..................................................................................................... 41

7 Discussion .................................................................................................................... 43

7.1 Mineral Resource Evaluation.......................................................................................... 43

8 Conclusion ................................................................................................................... 50

9 Recommendations ....................................................................................................... 52

10 References ................................................................................................................... 53

APPENDIX 1 ......................................................................................................................... 55

Sample point location maps for Tailings Stockpiles .................................................................. 55

APPENDIX II ......................................................................................................................... 62

Table showing sample point data ............................................................................................. 62

APPENDIX III ........................................................................................................................ 73

Gold Distribution Models of Tailings ......................................................................................... 73

APPENDIX IV ....................................................................................................................... 81

Operation profitability Charts ................................................................................................... 81

APPENDIX V ........................................................................................................................ 88

Definition of Terms .................................................................................................................... 88

Inferred, Indicated and Measured Mineral Resource according to NI43-101 .......................... 89




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LIST OF FIGURES

Figure 1: Drainage Au Summary & STRM topography. Map produced by Omai Gold Mine

Limited ............................................................................................................................................ 9

Figure 2: Auger, Banka, Rock & Production Drill hole Summary .................................................. 10

Figure 3: BTA Statistical Summary ................................................................................................ 26

Figure 4: BTD Statistical Summary ................................................................................................ 27

Figure 5: BTF Statistical Summary ................................................................................................. 28

Figure 6: BTG Statistical Summary ................................................................................................ 29

Figure 7: BTH Statistical Summary ................................................................................................ 30

Figure 8: BTI Statistical Summary ................................................................................................. 31

Figure 9: BTN Statistical Summary ................................................................................................ 32

Figure 10: Pie Chart showing Gold Distribution............................................................................ 36

Figure 11: Diagram illustrating the suggested circuit for reprocessing the tailings ..................... 40

Figure 12: Graph illustrating the connection between Operation Efficiency and Operating Grade

....................................................................................................................................................... 46

Figure 13: Profitability chart for operation at gold price $1200 USD/Oz ..................................... 48

Figure 14: Sample point layout for tailing BTA ............................................................................. 56

Figure 15:Sample point layout for tailing BTB .............................................................................. 56

Figure 16: Sample point layout for tailing BTC ............................................................................. 57

Figure 17: Sample point layout for tailing BTD ............................................................................. 58

Figure 18: Sample point layout for tailing BTE ............................................................................. 58

Figure 19: Sample point layout for tailing BTF.............................................................................. 59

Figure 20: Sample point layout for tailing BTG ............................................................................. 59

Figure 21: Sample point layout for tailing BTH ............................................................................. 60

Figure 22: Sample point layout for tailing BTI .............................................................................. 61

Figure 23: Sample point layout for tailing BTN ............................................................................. 61

Figure 24: Gold Distribution Model for BTA ................................................................................. 74

Figure 25: Gold Distribution Model for BTD ................................................................................. 75




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Figure 26: Gold Distribution Model for BTF .................................................................................. 76

Figure 27: Gold Distribution Model for BTG ................................................................................. 77

Figure 28: Gold Distribution Model for BTH ................................................................................. 78

Figure 29: Gold Distribution Model for BTI................................................................................... 79

Figure 30: Gold Distribution Model for BTN ................................................................................. 80

Figure 31: Profitability Chart for the operation at 40% efficiency for gold at $1200 USD/Oz. .... 82






LIST OF MAPS

Map 1: Omai Location Map .......................................................................................................... 13

Map 2: Location of Project Area ................................................................................................... 14

Map 3: Geological Summary of Omai. Map produced by Omai Gold Mines ............................... 16

LIST OF TABLES

Table 1: Sample grid Configurations ............................................................................................. 19

Table 2: Summary of Average Au Grades for Tailings .................................................................. 25

Table 3: Tailing Mineral Resource Estimate Summary ................................................................. 35

Table 4: Tabulation of cost associated with elements of the operation ...................................... 41

Table 5: Monetary value of total Recoverable Gold in Tailing ..................................................... 43

Table 6: Table showing the monthly requirement of the Operation ........................................... 45

Table 7: Summary of Sample Data ................................................................................................ 63




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1 INTRODUCTION

Metallica Commodities Corp. (henceforth the firm/company) is a locally based, foreign funded

mining company operating in Guyana under license, in the extraction and recovery of gold and

other precious metals. The company first began its operation targeting, like most local miners

alluvial gold leads, but had embedded in its future developmental plans the intent of starting a

project aimed at reprocessing the tailing product from gold mining activities, specifically those

produced by hydraulicking (sluicing). The firm has since acquired a section of land within the

West Omai PL of which a portion has been almost entirely worked by artisanal miners. The

miners employed hydraulicking as their method of extraction and subsequently left a number

of tailings dumps dispersed across the region. It is the opinions of the firm that these tailings

may contain gold in amounts worthy of mining. As such, Metallica CC has opted to conduct a

tailing evaluation program that will aid the decision making process leading up to the possible

reprocessing of the tailings.

1.1 Background

Gold mining has been conducted in Guyana for over 50 years with millions of ounces being

produced over this period. Much of this production is credited to the many artisanal mining

operations within the country’s interior lands. With the exception of the pass 20 years, gold was

traditionally and still is in some cases, produced through basic hydraulicking. Although this

method has been responsible for much of the gold produced in Guyana on the small and

medium scales, it is well known that the efficiency of this method is marginal. The method is

largely applied to alluvial gold deposits where gold is hosted in slightly compacted soil.

Hydraulicking entails the use of high pressure water jets to disintegrate the soil to form slurry, a

mixture of the soil and water. Using gravel pumps, the slurry is pumped up gradient to a

mounted apparatus called a sluice box. The slurry is then allowed to run over the sluice where

gold, being denser than the other constituent settles in the undercurrent of the flow. The base

of the sluice is covered with mat like material that can trap the gold traveling along it in the

undercurrent. Mercury is usually added before and/or after this process for amalgamation.




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Gold trapped in the mat is freed through a process of beating the mats into a basin filled with

water. After all the mats have been beaten, the water is removed from the basin leaving behind

the freed sediments. Panning and amalgamation is employed at this point to extract the gold

from this concentrate. Though this method of mining is largely profitable due to the low capital

cost as compared to other methods, it is still considered inefficient because of the high loss of

gold to the tailings. The overall recovery of makeshift sluice box ran by small-scale miner is

about 40-50% (BGS, 1997). Because of this fact, it is no surprise that some local miners engage

in reprocessing their own tailing and those of others to potentially recover lost values. These

efforts are usually still substandard due to the fact that miners utilize the same hydraulicking

method previously employed yielding only marginal recovery from the total potential resource.

Advances in the technology and increased knowledge in the area of mineral processing and

recovery have resulted in equipment and methods with higher efficiency. Their efficiency has

greatly surpassed those of ancient methods and has made it possible to recover values from

materials that were once considered low grade. Enhanced gravity equipment such as the

Knelson concentrator promises recovery of at least 80% in optimal configuration. Methods like

froth flotation go even further by offering recoveries of up to 99%. These resources coupled

with the present gold prices1 have sparked increasing interest into the reprocessing of tailings.

Through the period for which Metallica CC. has been operating in Guyana, efforts were put into

obtaining the necessary equipment, technology and personnel required for effective and

efficient mining practices. These efforts resulted in the acquisition of high efficiency processing

units such as trommel screens, Icon and Knelson gold concentrators and shaking table. With

these resources at hand, the company is ready to expand its operation to include a tailing

reprocessing program.

This opportunity was presented following investigation of a section of land within the

relinquished west Omai PL. The Bangalee Creek area is located within this PL and was for the

period February 1997 – February 2005 under investigation by Omai Gold Mine Limited. Omai

Upon completion, Omai Gold Mines Limited subsequently produced a report detailing this work

and highlighted within it the potential of the area to host a sizable alluvial deposit for small

1 As of July, 2012 world gold price taken from www.goldprice.com was averaged at USD $1600.




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scale mining. This knowledge led to the investigation of the area, upon which it was discovered

the land was already worked by artisanal miner. Littered with tailing dumps the firm decided to

conduct an evaluation of these tailing since it is of the opinion the tailings may contain gold in

amounts worthy of mining.

The aim of this project is to evaluate the tailings dumps in the Bangalee creek area to

quantitatively assess the presence of gold following which the preliminary feasibility of mining

the tailings will be assessed. This report documents the objectives, methodology and findings of

the project.




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2 OBJECTIVES

I. To delineate and quantify the occurrence of gold within tailings dumps of the Bangalee

Creek Area, West Omai

II. To conduct a feasibility for a tailing reprocessing operation of the tailings material in the

Bangalee Creek Area, West Omai




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3 LITERATURE REVIEW

3.1 Previous Work

3.1.1 Omai Gold Mines Limited – Final Report of West Omai

Through the period February 1997 to February 2005, Omai Gold Mines Limited had conducted

exploration work on the Western part of the Omai block and subsequently produced a report

detailing the activities and findings of the project. Omai Gold Mines undertook work in the area

to investigate the possible westward extension of the Wenot contact into this section of the PL.

Though the Wenot contact was successfully traced westward, then along a cross fault to the

south into the South Boundary Target area and drilling proved that it was still locally

mineralized, grades and volumes were very small. Work done by Omai Gold Mines Ltd included

drainage sampling (some multi-element), geological mapping, grid-based auger sampling,

Banka, Production and Diamond Drilling. This research paper bears reference to the

aforementioned report because it presents currently the primary, most recent and only

documented source of previous work data for the Bangalee Creek Area.

Finding: Bangalee Creek Alluvial Target

Primary mineralization occurs in irregular shear related quartz veinlet zones on the northern

flank of Quartz Hill. This may represent a splay off the Wenot Contact. Primary mineralization in

the Quartz Hill area has been extensively eroded to form the alluvial/colluvial resource in

Bangalee Creek.

This alluvial area extends downstream into the western part of the Omai PL. As shown in Figure

2, locally good (>2ppm Au) auger or Banka drill results occur in the low-lying parts of the

Bangalee Creek area, related to a similar alluvial zone. Gold was found to be restricted to a thin

gravel layer lying below 2-3m of barren sand, and in the western part of the area to supergene-

enriched laterite zones max 6m thick underlain by barren saprolite.

The laterite area was drilled, in the hope that gold was derived from a mineralized structure

hidden under the alluvium; drilling also led to the conclusion that the quartz float, soil




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anomalies and possibly the Bangalee Creek mineralization were all mudslide-type occurrences,

with a source(s) uphill to the south.

3.1.2 Historic Mining Activities

As was previously mentioned, the present tailing stockpiles existing throughout the Bangalee

Creek area were the product of previous hydraulicking activities. Though the legitimacy of the

data is questionable, conversations with locals within the vicinity of the study area have

indicated that production values obtained from single operations averaged 30 ounces per

operation cycle (4-5 days).




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Figure 1: Drainage Au Summary & STRM topography. Map produced by Omai Gold Mine Limited




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Figure 2: Auger, Banka, Rock & Production Drill hole Summary




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3.2 Review of other related literature

Examples of exploration work conducted on tailings were sought and revised from many

source. Particularly relevant to this project proposal were the methods used to gather, analyze

and present the various data sets. Stated below are examples from some of the methodologies

revised.

3.2.1 Sona Resources Corp – Blackdome Mine Tailings, British Colombia

Type and Extent of Drilling

Prior to drilling, a grid was established over the entire tailings pond. A baseline trending 065°

was established along the crest of the tailings pond. A series of 10 lines were run perpendicular

to the baseline. Lines were spaced at 15 meter intervals except at the boundaries of the tailings

pond where a reduced spacing of 10 meters was used. Drill holes were spaced at 30 meters

except near the south end and margins of the pond where shorter intervals were necessary.

Sampling Methods

The tailings pond was sampled with vertical holes drilled along at 30 meter intervals along lines

spaced at 15 meters or less. Samples were collected at five-foot (1.5 m) intervals throughout

the holes. Not all the holes could reach the bottom of the tailings pond. This was the case for

holes close to the dam and was due the thickness but also the type of material in this area

(sandy fill).

3.2.2 Lake Victoria Mining Company, Singida, Tanzania

This company was actively engaged in a tailing sampling and evaluation project. The project

targeted tailing piles of small scale artisanal miners at the Kilimanjaro Mining Company’s

Signida “Gold Rush” area. The tailing piles are a result of the past five years of small scale

mining, processing and gold recovery operations by the miners. The sampling method utilized in

this project provided some insight into the design of the sampling program for the current

research. At Lake Victoria Mining Company, sampling was conducted on a grid spacing of 4

meters by 4 meters using augers. The holes were augered with samples taken at each one

meter interval.




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4 SCOPE OF WORK

This report outlines the activities relating to and the subsequent findings of an exploration

program designed and executed for the monetary evaluation of gold tailings. The project was

conducted under Metallica Commodities Corp, for whom the findings of the project best

served. The completion of the project and compilation of this report relied on the following

activities:

a. Location and Identification of gold tailings within the area of study

b. Sampling and laboratory analysis of those samples for data acquisition

c. Interpretation and analysis of data for modeling and tailing resource estimation

d. Operational cost analysis for the proposed tailing recovery program

e. Deduction of preliminary project feasibility based on proposed operation cost and

estimated resource value.

4.1 Location and Access

The area of study is located within a section of land owned by the company in the relinquished

West Omai PL. The project area sits on a section of land in the vicinity of the Bangalee Creek

which is located on the right bank of the Omai River (Map 2). The Omai PL is located within the

Potaro Mining District; Central Guyana, and is approximately 165 Km South-southwest of

Georgetown (Map 1).

The Omai area is readily accessible via the Linden/Lethem road. It requires an approximate 60

miles of driving along this main road before branching off right onto the Omai road. The Omai

road leads to the Omai crossing which takes you across the Essequibo River. Once on the right

bank of the river, another 10 minutes driving along the Omai road will give you access to the

Omai area. The sometimes potholed Linden/Lethem road can lend for an uncomfortable drive

but usually an easy one. The Omai PL can also be accessed by plane landing at the Omai airstrip.




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Map 1: Omai Location Map




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Map 2: Location of Project Area




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4.2 Geological Setting

This project seeks to evaluate tailings material and as of such, the regional geology will not

explicitly represent the current nature of the resource assessed. It however provides some

background information.

4.2.1 Regional Geology

The Bangalee Creek and its environs are located in the Paleoproterozoic Barama-Mazaruni

Supergroup, a greenstone terrain deformed and metamorphosed during the Trans-Amazonian

orogeny, a tectono-magmatic event bracketed between approximately 2.3 and 1.9 Ga. The

greenstone belt sequence comprises alternating felsic to mafic and ultramafic volcanic flows

interlayered with thick sedimentary units2. The Barama-Mazaruni Supergroup is intruded by

acid dykes (e.g. quartz feldspar porphyry) and small granitic intrusions. Larger intrusions of

Younger Granite locally result in contact metamorphism, resulting in the formation of

amphibolites.

4.2.2 Local Geology

The Bangalee Creek tailings are located within the creek’s valley. The surrounding laterite hills

are capped with unconsolidated White Sand with thickness of up to 10m. Primary

mineralization in the Quartz Hill area has been extensively eroded to form the alluvial / colluvial

resource in Bangalee Creek. The Tailings occur disperse through the creek valley for a distance

of about 1km. The tailing comprise of a partially compacted gravel-sand matrix usually

underlain by a moderately plastic silt-clay horizon.

2 Gibbs and Barron, 1993




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Map 3: Geological Summary of Omai. Map produced by Omai Gold Mines




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5 METHODOLOGY AND DESIGN APPROACH

The following section describes the procedures used in the execution of this project. The main

activities and task are here described under the following sections; Data Acquisition, Data

Analysis and Evaluation.

Objective 1

5.1 Data Acquisition

5.1.1 Desk Study

Consultation of various literatures and other relevant sources of information on the

geographical project area and the topic of tailings evaluation was essentially the first step in

execution of the project. Articles relating to tailings evaluation programs were largely sourced

from the internet. Such documents provided information on the various concepts, strategies,

methodologies, standards and resources considered and utilized globally for projects aimed at

reprocessing gold bearing tailing dumps/stockpiles. These documents were particularly helpful

in the development of the sampling plan used in this project. Such documents are cited in the

literature review section of this report and thus will not be expounded on here. The Guyana

Geology and Mines Commission (GGMC) served as the major source of information relevant to

the project area i.e. the Bangalee Creek area. Reports such as the West Omai Final report cited

in the literature review provided useful background information into the region. The GGMC also

served as a source for maps of the area. Discussions with locals provided useful background

information on the region of study.

5.1.2 Site Characterization

Field wok commenced with activities related to site characterization. A planned 3 days trip was

used to observe and describe the Bangalee Creek area and tailings dumps which it contained.

Activities conducted during this period included identification of tailings dumps, determination

of accessibility to individual tailings and describing and measuring of accessed dumps. The

locations of accessed dumps were marked and recorded using the Garmin CSX 60 GPS. Each




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tailings dump was given a field name for identification which was entered into a field log book

along with description, measurement and location data. This information (except description)

was also placed onto a strip of flagging tape tied to a stick and stuck upright on each tailing.

This preliminary site investigation work along with knowledge garnered from consulted

documents provided the basis for designing a comprehensive and strategic sampling program.

From the preliminary field work conducted, it was seen that the tailings dumps varied in their

shape, size and material distribution. These differences were quickly identified as a reason for a

sampling method that would be adaptive to the characteristics of each tailing. The program

would need to, at the same time, be able to sample each tailing adequately without producing

excessive amounts of samples.

5.1.3 Sampling Strategy and Technique

The sampling strategy employed was aimed at providing detailed information on the

characteristics of the tailings dumps. As previously mentioned the variations in shape, size and

depth of tailings dumps, required the sampling method to be adaptive to the various tailings

dumps for adequate representation of each. Sampling was done using regular grids, where

samples were collected at grid line intersections occurring on the tailings. Grid configuration

varied based on the calculated rectangular area of the individual tailings (see Table 1). In

establishing the grid, the first point was positioned at the apex of the dump from which the grid

developed outwards to the extent of the tailing.

This style of grid and sampling was selected because it ensures maximum cover of the tailings

dump while maintaining an adequate sample density. The utilization of a single fixed grid across

all tailings would have produced one of two possible costly scenarios. In the first scenario

where a standard small grid is utilized (e.g. 5m×5m), in large tailings this would have produced

a large number of sample points, resulting in the over sampling of the mass. Likewise, a

significantly larger sample output for laboratory analysis would have been produced resulting in

longer processing and overall increase in cost. For the second scenario where a standard large

grid is utilized (e.g. 7m×7m), small tailings stockpiles would have been inadequately sampled

and therefore inadequately assessed resulting in the underestimation of the potential mineral

resource. The larger sample spacing on smaller tailings would have reduced the number of




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sample points on these tailings, which could be costly since critical point may be missed due to

wide spacing.

Table 1: Sample grid Configurations

Description Grid Configuration

Tailing having dimensions less than 144m2 4m×4m

Tailing having dimensions between 144m2 – 400m2 5m×5m

Tailing having dimensions between 400m2- 900m2 6m×6m

Tailing having dimensions greater than 900m2 7m×7m

The distribution of unrecovered gold values in these tailings stock is a process controlled by

factors ranging from particle size and density to discharge rate from the sluice. However, it is

know that from time to time, these miners would employ various means, physical or

mechanical, of repositioning the discharge from the sluice in an effort to keep the discharge

point clear from material piling up and “choking” the system. As of such, this program was

designed without the assumption that the distribution of the mineral of interest was not

exclusively controlled by the normal dynamic factors.

Samples were collected at 1m depth interval (where applicable) at sample points established

from the grid configuration utilized. A motorized auger was use to drill vertically to the base of

the tailing at each sample point. The base of the tailing was interpreted as that part where a

very plastic “slushy” silt-clay material was intersected or where consolidated humic soil was

encountered. At each sample point, a sample was taken at each 1 meter interval of depth. The

material was retrieved from the auger bit and placed onto a plastic bag or into a bucket where

it was homogenized after which a sample of approximately 300 g was taken. Sampled materials

were placed into draw-string plastic sample bags. Sample ID, sample description, sample depth

interval and tailing location were recorded in sample books. Sample books consisted of a

perforated sheet that can be torn off and placed into the sample bags. The main segment in the

book and the perforated segment had matching sequence numbers, so samples can be

identified or correlated with filed description data. Upon completion of a sample point all tools




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and materials used during the auger and sampling process were rinsed thoroughly before being

used at the next sample point, this reduced the possibility of cross contamination.

Samples were transported from field to the laboratory in buckets in an effort to avoid damage

to the samples and spillage of sample content. Stacking of samples was avoided since this could

force water and suspended particles out of underlying samples, which in turn can contaminate

surrounding samples.

5.1.4 Sample Preparation and Laboratory Analysis

Samples collected in the field were subsequently submitted to the onsite Metallica CC

laboratory (previously Omai Gold Mines Laboratory) for preparation and analysis. The sample

preparation procedure was as follows:

1. Sample tags were removed and samples were placed into numbered pan.

2. Sample tag number and the number of the pan in which it was placed were recorded

into lab entry book for tracking.

3. Pans were then placed into a countertop hot air convection oven for drying at 150°C for

approximate 8-12 hrs.

4. After drying, each sample was passed through a riffle box to obtain two reduced splits.

5. The splits were placed into individual sample bags and labeled according to the sample

tag.

6. One split was sent for continued preparation while the second was placed into storage

as a duplicate sample.

7. The reduced split was sent for pulverization for size reduction and screening so that 85%

of the material passes the US 200 mesh.

8. From the pulverized product, 30g was extracted from each sample for fire assay.

Between each sample, laboratory equipment was cleaned to reduce the possibly of

contamination.




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5.2 Data Analysis

5.2.1 Processing and Analysis of Downhole Data

This process was achieved through the use of the geological modeling software, RockWork15.

All values obtained from the laboratory were entered into an Excel spread sheet along with

field data pertaining to each sample i.e. - field designation, sample location, sample description,

etc., these data were then imported into RockWorks15. RockWorks15 is an integrated software

package that can be used for the management, analysis and visualization of geological data. All

geographical and geologic data related to the samples obtained were managed and

subsequently processed utilizing the RockWorks15 software package.

5.2.1.1 Mapping

The first step required for the analysis of downhole data is the 2-Dimensional surface projection

of the drill/auger holes. This was done for each tailing analyzed utilizing the grid array that was

assigned to the tailing. This produced a surface map showing the relative location of the sample

points on the tailing, (see Appendix I). Entering the grid based sample location data was easily

achieved in the RockWorks15 software since it provided an excel spreadsheet type utility for

entering this data. This data was entered in a Cartesian coordinate format, where the locations

of sample points were given relative to a central point/origin and two perpendicular axes.

Associated data entered included sample ID and total depth value for each sample point.

Though not necessary for the mapping of the sample point, the hole depth was required for the

next step; 3D Modeling.

5.2.1.2 Grade Distribution Models

Following the input of the location data for all sample points on a given tailing, the next step

involved the entering of the down hole data related to the sample points. For the purposes of

the project, this was mainly the sample assay values. Again, this data was easily entered into

the software through its downhole management tool. The assay value and the depth interval

for the specific sample point which they represent were entered into downhole management

tool. This was done until each sample for the specific tailing dump was accounted for. The

proper entry of these data allowed for the production of 3D grade distribution models.




Page | 22

October, 2012

Appendix III illustrates the grade distribution models for the tailings. Note that the outlines of

the models do not necessarily represent the actual shape of the tailings bodies.

5.2.1.2.1 Compositing

Length weighted compositing of assay values for individual sample points was done prior to the

generation of models and calculation of average tailing grade. This task was accomplished in

RockWorks15. Length weighted compositing refers to the process in which the assay values for

a specific drill hole are multiplied by their corresponding thickness. These values are then

summed and divided by the sum of their thickness, thus giving an average value for the drill

entire section or interval. This was done to obtain an average grade for each sample point on

the tailing, since each point comprised of a number of downhole samples at various depths.

Compositing was done for the entire length of the drill hole rather than in increments.

Compositing of downhole data is usually a very useful tool since it effectively reduces the

number of samples to be processed for modeling and other computations.

5.2.1.2.2 Modeling Algorithm

Modeling the gold grade distribution for the individual tailings was done using the inverse

distance anisotropic solid modeling method. This method essentially utilizes the inverse

distance algorithm. With this algorithm, a voxel point value is assigned based on the weighted

average of neighboring data points, and the value of each data point is weighted according the

inverse of its distance from the voxel node, taken to a power (an exponent of “2? = Inverse-

Distance squared, “3? = Inverse-Distance cubed, etc.). In assigning a voxel node value, the

anisotropic method used the closest point within each 90 degree sector around the voxel. The

anisotropic method was chosen based on the nature of the assay data. The method reduced

the bull’s-eye effect that the raw data presented and gave a better representation of grade

distribution.

5.2.1.3 Statistical Analysis

While grade distribution models are useful for evaluating and visualizing the distribution of

sample assay data, mineral resource estimation calculations are dependent on the




Page | 23

October, 2012

development of an average grade for the deposit being considered. The average grade for the

various tailings investigated was computed using the RockWork15 software. Based on the input

data, the program calculated parameters such as the mean, standard deviation; variance and

coefficient of variation for the data set for each tailing dump (see pg.26-32). The mean was

used as the average grade value for the entire tailing body. A summary of the computed

average Au grades are provided in Table 2, pg.25

Object 2

5.3 Evaluation

Findings from the data analysis were incorporated into this phase which sleeked to determine

whether the feasibility of reprocessing the tailing dumps. Tailings grade values obtained from

the previous phase were applied to material tonnage data gathered to further define the

resource and estimate potential return based on current gold prices. Operational cost for the

reprocessing of the tailings was estimated based on the present operational cost incurred by

the firm. Profitability of the proposed operation was evaluated from this data.




Page | 24

October, 2012

6 RESULTS AND ANALYSIS

6.1 Calculation of Tailing Average Au Grade

The averaging of the Au grade for each tailing was done utilizing the composited sample point

values. Because compositing produced an average value for the entire length of an auger hole,

it made it simpler to compute the average tailing grade since a smaller population can be

considered. The value of each sample point was taken to represent the average grade for a

block/square enclosing that point. The size of the block was determined by the grid

configuration utilized to sample the tailing. For example, if a 5m × 5m grid configuration was

used to sample the a tailing dump, then the value of each sample point on that tailing would

represent a block of 5m × 5m, like wise for a 6m × 6m grid configuration, each sample point

would represent a block 6m × 6m in dimension.

This method of analysis raises the question of area of influence of individual sample points.

Based on this methodology, it means that for larger tailing it is expected that the value of a

single sample point is expected to extent or influence possibly 2 times the area of a similar

sample point on a smaller tailing. It however is the opinion of the author that this does not raise

an issue with the computation of the overall grade, since it was accommodated for in the

planning stage of this project where a maximum grid configuration and by extension, area of

influence of 7m × 7m will be used. The average Au grades were calculated as the sum of the

individual sample grades for a tailing divided by the sample population being observed. Since

the composited data was used, the population size was always equal to the number of sample

points on the tailing.




Page | 25

October, 2012

Table 2: Summary of Average Au Grades for Tailings

Field Designation Tailing Coordinate (PSAD 56) Au Grade (g/t)

BTA 21 N 0302671 0602883 0.708

BTB 21 N 0302646 0602883 0

BTC 21 N 0302722 0602871 0

BTD 21 N 0302733 0602818 0.499

BTE 21 N 0302702 0602818 0

BTF 21 N 0302655 0602840 0.44

BTG 21 N 0302619 0602807 0.24

BTH 21 N 0302692 0602719 0.463

BTI 21 N 0302665 0602701 0.234

BTN 21 N 0302667 0602739 1.22




Page | 26

October, 2012

Figure 3: BTA Statistical Summary




Page | 27

October, 2012

Figure 4: BTD Statistical Summary




Page | 28

October, 2012

Figure 5: BTF Statistical Summary




Page | 29

October, 2012

Figure 6: BTG Statistical Summary




Page | 30

October, 2012

Figure 7: BTH Statistical Summary




Page | 31

October, 2012

Figure 8: BTI Statistical Summary




Page | 32

October, 2012

Figure 9: BTN Statistical Summary




Page | 33

October, 2012

6.2 Mineral Resource Estimation

Much of the work and analysis thus far were part of a process aimed at providing information

regarding the quantity of the mineral of interest contained within the depositional bodies

investigated. In order to quantify the indicated mineral resource hosted in the tailings dumps,

two parameters had to be first obtained, the average gold grade of the tailing and the material

tonnage of the tailings dump. The average Au values for each tailing and the method of

obtaining those values were previously discussed in section 6.1 and thus will not be revisited

here. The material tonnage of each tailings dump was obtained through volume × material

density calculation. The volume was obtained by multiplying the calculated area of the tailings

dump, i.e. measured length by width, by its average depth which was taken as the quotient of

the sum of sample hole depths in a tailing divided by the number of sample points. The process

of determining the mineral resource of a tailings dump is shown below.

Calculations 1

a. Calculation of Material Tonnage for BTA

Length of Tailings Dump = 22m Material Density = 1.3 t/m3

Width of Tailings Dump = 16m Average Depth of Tailings Dump = 1 m

i. Area of Tailing = Length x width

= 22m 16n

= 286m3

ii. Volume of tailing = Area of Tailing x Depth of Tailing

= 286 m2 × 1 m

= 286 m3

iii. Material Tonnage of Tailing = Volume of Tailing × Density of tailing material

= 286 m3 × 1.3t/m3

= 371.8 Tons




Page | 34

October, 2012

b. Calculation of mineral resource estimate for BTA

Material tonnage of Tailings = 371.8 tons

Average Au grade of tailing = 0.708 g/t

i. Total gold resource of BTA = Material tonnage of tailing × Average Au grade off tailing

= 371.8 tons × 0.708 g/t

= 263.23 g

Unit conversion factor

1 Ounce of Au = 31.1 grams

So that 263.23 g of Au = 263.23g / 31.1 �

��

= 8.46 Oz

This method of estimating the gold contained within the tailings material was repeated for each

tailing dump investigated. The actual process of calculating and tabulating these values was

done using the Microsoft Excel component of the Microsoft Office software package. A

summary of these calculations is shown in Table 3: Tailing Mineral Resource Estimate Summary,

along with the total material and mineral resources.

The mineral resource estimation process was performed in such a manner that all tailings

dumps reporting an overall grade of zero grams per ton of material were excluded from the

tabulation of the total material tonnage. This was achieved by using a zero value for the

material density for those tailings. What this produced was the total tonnage of gold bearing

tailings within the area. Using that value along with the total gold value, it was possible to

calculate an average grade for all mineral bearing tailings bodies.




Page | 35

October, 2012

Table 3: Tailing Mineral Resource Estimate Summary

Tailing

Dump

Field

Desig.

Tailing

Coordinate

(PSAD 56)

Width

(m)

Length

(m)

Area

(sq. m)

Avg.

Depth

(m)

Volume

(Cu. m)

Material

Density

(Ton/Cu

. m)

Material

Tonnage

(Tons)

Au

Grade

(g/t)

Au in

tailing

(g)

Au in

tailing

(Oz.)

1 BTA 21 N 0302671

0602883

13 22 286 1 286 1.3 371.8 0.78 263.23 8.46

2 BTB 21 N 0302646

0602883

12.5 26.5 331.25 1 331.25 0 0 0 0 0

3 BTC 21 N 0302722

0602871

11 18 198 0.5 99 0 0 0 0 0

4 BTD 21 N 0302733

0602818

10 14.5 145 1.3 188.5 1.3 245.05 0.499 122.28 3.93

5 BTE 21 N 0302702

0602818

18 13 234 2 468 0 0 0 0 0

6 BTF 21 N 0302655

0602840

18.5 30.5 564.25 1.7 959.23 1.3 1246.99 0.44 548.67 17.64

7 BTG 21 N 0302619

0602807

15 19 285 1.5 427.5 1.3 555.75 0.24 133.38 4.29

8 BTH 21 N 0302692

0602719

10.5 26.5 278.25 1.7 473.03 1.3 614.93 0.463 284.71 9.15

9 BTI 21 N 0302665

0602701

24 30.5 732 1.7 1244.4 1.3 1617.72 0.234 378.54 12.17

10 BTN 21 N 0302667

0602739

22 27 594 1.5 891 1.3 1158.3 1.22 1413.12 45.44

Total

5810.545

3143.957

101.09



October, 2012

Figure 10: Pie Chart showing Gold Distribution

12%

45%

Percentage Distribution of Gold across the Tailings Dumps


8%0%

0%

4%

18%

4%

9%


Page | 36

0%

18%


BTA

BTB

BTC

BTD

BTE

BTF

BTG

BTH

BTI

BTN




Page | 37

October, 2012

Calculations 2

a. Calculation of Average grade of Tailings

Total tonnage of mineral bearing tailings material = 5810.54 tons

Total gold in mineral bearing tailings = 3143.96 grams

i. Average grade of tailings = ��

��

= ��.��

��.��

= 0.541 g/t

By establishing the overall average grade of the mineralized tailings, it becomes simpler to

assess the resource and determine its economic viability. The average grade of the tailings

material will be used for evaluating the tailings to determine whether the current resource is

substantial enough for reprocessing or if, should it be necessary, additional tailings dumps

should be eliminated from the population in order to produce a higher grade of material for

mining.




Page | 38

October, 2012

6.3 Operational Cost Estimation

Exploratory work and computer aided analysis of the data it provided have resulted in the

establishment of an “indicated mineral resource” of 3143.97g (101.09 Oz). of gold contained

within 5810.54 tons of tailings material. The stated mineral resource estimate is the product of

the ten tailings dumps within the Bangalee Creek area targeted and investigated by this project.

The indicated tonnage of tailings material for which the mineral resource is contained, was

deduced following the removal of tailing reporting a grade value of zero. Based on these

figures, it has been determined that the total average grade of the mineralized tailings dumps is

0.541g Au/t.

For the purposes of this report, a brief description of the firm’s proposed mining method is

presented in the following subsection.

6.3.1 Mining Method

The circuitry of the main operations and processes is illustrated in Figure 11 and are here

expounded upon. Tailing material selected for reprocessing would be excavated and heaped,

when possible, into one stockpile at a designated point for processing. High pressure water jets

will be used to disintegrate the material in the stockpile and create slurry. The slurry under the

influence of gravity will move down an established inclined trench in to a small sump. Using a

gravel pump, the slurry will be sucked up and sent up slope to a trommel for sieving. The

trommel, a cylindrical rotating sieve, will separate the passing material in to two size fraction:

+2mm and -2mm respectively. The +2mm material produced at opening at the other end of the

trommel will be stocked and stored for backfilling of pits within the area. Material reporting to

the undersize of the 2mm mesh will pass through the base of the trommel and down a trough

which will direct the material to a second, larger sump. Material accumulated in this sump will

be sent, via a gravel pump, to a Knelson Concentrator for processing. The Knelson, which is a

enhance gravity concentrator; will cause the gold and other dense particle to lodge within the

riffle in its bowl while expelling the lighter presumably waste particles. The waste materials are

channeled away from the Knelson, via a series of pipes, and into a third sump for storing, where

it can be excavated and used for backfilling of surrounding old pits and ponds. After an




Page | 39

October, 2012

estimated run time of about 1 hour, the Knelson will be stopped to retrieve the concentrate.

Subsequently the concentrate will be sent for further processing on the shaking table to

recover the gold. The operation is expected to require 9 persons:

• (1) Excavator Operator

• (1) Team Leader

• (3) Jet men

• (2) Morack Men

• (2) Pitmen

Based on the previous data provided by Metallica CC Guyana Inc., an average processing rate of

90ton solid/day was calculated for its operations utilizing this circuit and method of mining.

Based on specifications provided for equipment being utilized within the circuit, the operation

should be capable of processing at minimum 20 ton solid /hour. Processing time is expected to

be about 8-9 hr. per day. This translates to an average processing rate of 170 ton of solid per

day of operation, a figure that greatly surpasses the current average. It is clear from this

comparison that the current operation is functioning far below capacity is thus inefficient.

Though it is currently unknown, the percentage mineral recovery of the operation is capable of

exceeding 80% efficiency at optimal performance. The Knelson concentrator, the primary

concentrator in the below circuit, at optimal setting is capable of recovering 95% of all gold

particles presented to the equipment.




Page | 40

October, 2012

Figure 11: Diagram illustrating the suggested circuit for reprocessing the tailings

SUMP SUMP STOCKPILE

FEED TROMMEL

SHAKING TABLE

+2mm -2mm

GOLD

KNELSON

CONCENTRATOR

CONCENTRATE

Pumping

Pumping

SLURRY

TAILING

HOLDING

POND

BACKFILL




Page | 41

October, 2012

6.3.2 Cost Assessment

A projection of the operational cost was made based on the proposed mining method and data

provided by Metallica CC Guy Inc. The calculation was done for a 25 day working month as

suggested by the firm. The maximum incurrable cost was determined utilizing all available data.

This value was taken as the costs encountered for the firm to mine and process the tailings for a

full 25 days period. It is expected that preparatory and auxiliary work would require less

resources and as such would generate a smaller monthly cost. Table 4 shows the tabulation of

the projected maximum monthly cost.

Since all necessary equipment, parts and machinery for operation are already available; such

costs were not tabulated into the expense of the operation.

Table 4: Tabulation of cost associated with elements of the operation

Item Unit Cost Qty. Per Day Cost

(Total)

Monthly Cost

(25 days)

Excavator Operator 8200/day 1 8200 $ 205,000.00

Jet man 3500/day 3 10500 $ 262,500.00

Morack Man 3500/day 2 7000 $ 175,000.00

Pit Man 3500/day 2 7000 $ 175,000.00

Team Leader 4000/day 1 4000 $ 100,000.00

Fuel (Excavator) 48000/day 48000 $ 1,200,000.00

Fuel (Generator) 18800/day 18800 $ 470,000.00

Food and Accommodation 3200/per/day 9 28800 $ 720,000.00

Sub-Total $ 3,307,500.00

Miscellaneous 20% $ 661,500.00

Total GYD $ 3,969,000.00

@ 200 GYD/USD USD $ 19,845.00

@ 1200 USD/Oz Gold (Oz) 16.54

@ 1600 USD/Oz Gold (Oz) 12.40

As seen from the table above, the total maximum monthly cost incurrable to the firm is

approximately $3,969,000 Guyana dollars. Using a low gold price of $1200 USD/Oz and gold

high price of $1600/Oz for the period July, 2012 – December, 2012, the equivalent of the total




Page | 42

October, 2012

expense in gold was calculated. This figure in gold is between 12-17 ounces per month. This

means that in order for the operation to recover its monthly incurred expense, it must produce

12.40 or 16.54 ounce of gold at $1200 USD/Oz or $1600 USD/Oz respectively.

Since Metallica CC is currently involved in trucking of stones out of Omai, much of the necessary

facilities and therefore capital cost has been previously established and invested.




Page | 43

October, 2012

7 DISCUSSION

7.1 Mineral Resource Evaluation

The estimated quantity of the mineral of interest within the investigated bodies and the cost

associated with processing and recovering the mineral were obtained and discussed in previous

sections. It is now the aim of this section to assess this resource based on projected operational

cost to determine, and hence make preliminary pronouncements on, the economic viability of

reprocessing the investigated tailings for their contained mineral resource.

Through methods of compositing, modeling and statistical analysis of downhole data obtained

from samples taken from the 10 tailings dumps, an “indicated mineral resource” of 101.092 Oz

of gold was identified. Screening of tailings reporting an overall grade of zero resulted in a

reduction of the overall material tonnage being considered. Because the tailings eliminated

from the set had an average grade of zero, the initial calculated mineral resource value was still

the same. Using the reduced material tonnage and the computed mineral resource value which

was contained within it, an average grade was calculated for the tailings deposits. This value

was 0.541g Au/t while the reduced material tonnage value was 5810.545 tons. The estimated

cost of an operation targeting these tailings utilizing the mining method described in section

6.3.1 was calculated at $19,845.00 USD per operating month i.e. – 25 days, or $793.80 USD per

day. This cost appears relatively minute when the value of the mineral resource is considered as

shown in Table 5. It is however important at this stage that two aspects of the operation, i.e. –

efficiency and processing rate, be considered prior to making any declarations. The

performance of the processing circuit in terms of mineral recovery and processing rate are

important to the successful exploitation of the mineral resource.

Table 5: Monetary value of total Recoverable Gold in Tailing

Indicated Mineral Resource

(Oz.)

Gold Price

(USD/Oz)

Value of Gold Resource

(US Dollars)

101.092

Min $1200.00 $122,310.00

Max $1600.00 $161,747.20




Page | 44

October, 2012

These parameters are here evaluated in conjunction with computed mineral resource and

operational cost data to determine whether the proposed operation will be able to feasibly

mine the tailings dumps and at what performance level.

As stated in section 6.3.1, the processing plant at its current level of operation, works at an

average processing rate of 90 ton/day of solids. This translates to an average of 2250

ton/month of solids. Using this value along with the calculated gold production requirement

shown in Table 4 on pg. 41, the average gold grade required by the operation can be calculated.

Calculations 3

i. Monthly Material Tonnage Processed = 2250 Tons

Monthly Required Production @ 1200 USD/ Oz = 16.54 Oz

Required Grade = �� !��"��!#��@�%��&'(/��

��"��#��

= �*.��+,

%%��-./0

= 0.00735

= 0.00735 × 31.1g/Oz

= 0.228585 g/t

ii. Monthly Material Tonnage Processed = 2250 Tons

Monthly Required Production @ 1600 USD/ Oz = 12.40 Oz

Required Grade = �� !��"��!#��@�*��&'(/��

��"��#��

= �%.��+,

%%��-./0

= 0.00551 Oz/Ton

= 0.00551g × 31.1g/Oz

= 0.1714 g/t




Page | 45

October, 2012

Table 6: Table showing the monthly requirement of the Operation

Tons/ Day (Alluvial) 90

Daily Required Production @ 1200 USD/ Oz (Ozs) 0.66

Daily Required Production @ 1600 USD/ Oz (Ozs) 0.50

Tons/Month (Alluvial, 25 day) 2250

Monthly Required Production @ 1200 USD/ Oz (Ozs) 16.54

Monthly Required Production @ 1600 USD/ Oz (Ozs) 12.40

Monthly Required Grade @ 1200 USD/Oz (Oz/Ton) 0.0074

Monthly Required Grade @ 1600 USD/Oz (Oz/Ton) 0.0055

Monthly Required Grade @ 1200 USD/Oz (g/t) 0.23

Monthly Required Grade @ 1600 USD/Oz (g/t) 0.17

As summarized in the above table, the required grades at USD $1200/Oz and USD $1600/Oz are

0.228g/t and 0.171g/t respectively. The current resource has been observed as having an

overall grade of 0.541 g/t, a grade that surpasses the operational requirements. While this

remains the grade of the tailings bodies, it does not signal the grade at which the processing

plant would be operating. The operating grade of the processing plant is determined by the

percentage mineral recovery of the unit. A high mineral recovery percentage translates to a

high operating grade while a low recovery would represent a low operating grade.

The graph on pg. 46 indicates how changes in the mineral recovery percentage can affect the

operational grade of the processing plant and as of such the overall recovery of the operation.

The blue line indicates the operating grade at a specific efficiency level, while the red and green

lines show the grade required when the price of gold is USD $1200/Oz and USD $1600/Oz

respectively. From the chart, it can be seen that a recovery level of approximately 43% is

required to meet a breakeven at a low gold price of $1200 USD/Oz. At a high gold price of USD

$1600/Oz, the operation would require a lower recovery percentage, approximately 32%, to

balance its expenses. Note that the recovery levels discussed here and indicated by the chart

are based on the current average processing rate being observed by the firm, i.e. 90 ton solid

per day.




Page | 46

October, 2012

Figure 12: Graph illustrating the connection between Operation Efficiency and Operating Grade

0

0.05

0.1

0.15

0.2

0.25

0.3

0.35

0.4

0.45

0.5

0.55

0.6

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

GR

AD

E

Percentage Recovery

Operating Grade at % Recovery Required Grade @ 1200 USD/Oz (g/t) Required Grade @ 1600 USD/Oz (g/t)




Page | 47

October, 2012

As of such, they only suggest what efficiency levels are required by the operation to cover cost

if its processing rate remains the same. It also indicates that either a higher mineral recovery

level and/or processing rate are required to make the operation profitable. If the operation is

capable of a higher mineral recovery percentage than that required by the minimum gold price

estimate, it would be able to gain profitable returns from re-processing the tailings. Achieving

processing rates above 90 tons/day would also improve profitability of the operation since this

would reduce the amount of time required to exhaust the current resource, which should

theoretical result in a lower accumulated cost over time. Based on the material tonnage and

the current processing rate of the operation, it was calculated that the operation would require

approximately 65 days of processing in order to deplete the current resource.

At a daily operational cost of USD $793.80, 65 days of operation would run a cost of USD

$51,597.00. This is with the exclusion of the cost associated with the approximate 14 days of

preparatory work needed to setup the operation. In total, the operation would require about

80 days of work to setup and process the 5810.545 tons of tailings material necessary to

recover the contained gold values. The graphs in Figure 13 and Appendix IV show how, for the

operation functioning at specific efficiency levels utilizing minimum gold price estimate of

$1200 USD, the accumulated cost can affect profitable returns as time progresses. At each level

of recovery efficiency, there exists a fixed maximum amount of the resource than can be

recovered. Because of this, there is also a maximum return that can be generated from mining

the resource at a given time. However, as time progresses the incurrable cost of extracting the

mineral increases due to accumulation. As of such, the less time taken to mine the resource

the greater the profitable returns at that time period. At day 80 of operation, accumulated

operation cost is expected to be $63,504.00 USD. From the graph in Figure 13, it can be seen

that for the operation to span an 80 days period, a recovery percentage greater than or equal

to 60% is necessary to generate a profit. It is clear from this and other graphs in Appendix IV,

the less time taken to reprocess the tailings, the greater the profitable returns at the various

efficiency levels.




Page | 48

October, 2012

Figure 13: Profitability chart for operation at gold price $1200 USD/Oz

$0.00

$25,000.00

$50,000.00

$75,000.00

$100,000.00

5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80

US

Do

lla

rs

DAYS

Accum. Cost 40% Efficiency 50% Efficiency 60% Efficiency 70% Efficiency 80% Efficiency 90% Efficiency




Page | 49

October, 2012

In section 6.3.1 of this report, the potential operating capacity of the circuit being employed

was indicated based on specifications for the equipment being utilized. It is estimated, the plant

should be capable at minimum, to process 170 ton of solid/day at a recovery efficiency of 80%.

At a processing rate of 170 ton/day, it is calculated, the plant would require 34 days of

processing to exhaust the current resource. This is almost half the time required at the current

processing rate. In total, the operation would require just about 40 days to setup and mine the

tailings dumps. The operational cost for 40 days will sum to $31,752 USD, half of that required

at 90 tons/day.




Page | 50

October, 2012

8 CONCLUSION

This study was aimed at investigating and evaluating tailings dumps within the Bangalee Creek

Area to determine whether they contained a sufficient gold resource for extracting through

reprocessing of the tailings material. Following the objectives set out in this report, the

following conclusions were derived:

• Objective 1 and Objective 2

Ten tailings dumps were investigated for this research paper. Of the ten, seven were

considered mineralized based on fire assay analysis of tailings samples. Three Dimensional

grade distribution models of the tailings showed there was no significant trend in the zoning of

gold values within the tailings investigated. Samples detected of having gold values came from

various points and depth within the tailings. This could be due to the fact that miners generally

reposition the material from the tailings so as to keep it from building at the discharge point of

the sluice.

There exist an “indicated mineral resource” of 101.092 Oz of gold within 5810.545 tons of

tailings material spanning the seven tailings bodies.

• Objective 3

The investigated tailings stockpiles have the potential to be profitably reprocessed for

contained gold resource, but only at acceptable operational levels. While sufficient in a high

efficiency circuit, the current processing rate of the plant of 90 ton/day is considerably low and

only produces marginal benefits. A total of 80 would be required at this processing rate to

setup the operation and mine the tailings. It was projected that this lengthy period would

accumulate, based on a daily operational cost of USD $793.80, an expense of $63,504.00 USD.

Based on this figure it was determined that operation would need to be at least 60% efficient in

mineral recovery for it to be profitable. However should the plant be able to operate at its

proposed high minimum capacity of 170 ton/day, the accumulated cost would only be $31,752

USD after the required 40 days of operation. As is clearly illustrated in Appendix IV, the return




Page | 51

October, 2012

after 40 days of operation is much more than that after 80 days. Recovery levels of at least 80%

are inferred based on researched equipment specifications. Once achieved, this would

considerably boost the profitability of the project.




Page | 52

October, 2012

9 RECOMMENDATIONS

1. A micro balance should be obtained for the laboratory to be used for weighting of beads

following fire assay. Currently the lab operates with a semi microbalance which is only

capable of reporting values to a minimum of 0.01 g.

2. Implementation of other sample analysis methods, such as Atomic Absorption

spectroscopy for detection of values below the detection limits of the fire assay method.

3. Sieve and material density analysis should be conducted to better inform the mining

method and processing plant setup.

4. The current processing circuit should be revised in light of metallurgical analysis in an

effort to optimize performance. A better performing processing plant would encourage

increased profitable returns.

5. Continued sampling and analysis of the remaining tailings within the Bangalee creek

area to expand the current resource.




Page | 53

October, 2012

10 REFERENCES

Books

1. Gibbs, A.K, Barron, C.N., (1993). The Geology of the Guiana Shield. Oxford University

Press, New York. [Accessed 6 December, 2011]

PDF Documents

2. 2003, Abbreviated Preliminary Assessment: Longshot Mine and Millsite. [pdf] Available

at www.fs.usda.gov/Internet/FSE_DOCUMENTS/fsbdev3_035008.pdf, [Accessed 11

December, 2011]

3. Body, K. and Lomberg, K. 2008, Elsburg Tailing Dams Complex: Independent Competent

Person Report. [pdf], RGS Global Consulting Pty Ltd. [pdf] Available at:

www.drd.co.za/b/files/Elsburg_cpr_report20aug08.pdf, [Accessed 6 December, 2011]

4. Fire River Gold, 2010, Fire River Gold Corp. Tailings Sampling/ Drill Program Averages

7.6 g/t Gold at Nixon Fork Gold Mine, Alaska [press release], Available at

www.firerivergold.com/i/pdf/040610FAU.pdf, [Accessed 6 December, 2011]

5. Fire River Gold, 2010, Fire River Gold Announces Update on Tailings Preliminary

Economic Assessment, Nixon Fork Gold Mine, Alaska [press release], Available at

www.firerivergold.com/i/pdf/061410FAU.pdf, [14 December, 2011]

6. Grunewald, W. and Makepeace, D. 2010, Technical Report on the Blackdome Mine

Tailing. [pdf]. Available at www.sonaresources.com/_.../NI_43-

101_Blackdome_Report_F1.pdf, [Accessed 11 December,2011]

7. Heesterman, L. 2008, Final Exploration Report on the Western Part of the Omai Mining

License PL 01/1997 for the period February 14th

,1997 to February 13th

, 2005. [pdf], Omai

Gold Mines Limited. [Accessed 27 July,2011]




Page | 54

October, 2012

8. Kuchling, K., Adams, P., et al., (2011). Technical report on the Las Lagunas Gold Tailings

Project. [pdf], kj Kuchling Consulting Ltd, Available at

http://panterragold.com/userfiles/file/NI43-

101%20Technical%20Report%2031_12_2011/PanTerra%2043-

101%20Las%20Lagunas%2043-101%20Tech%20Report%20Dec%2031%202011.pdf, [14

December, 2011]

9. Lake Victoria Mining Company Inc., 2009, Lake Victoria Mining Company’s Gold Tailing

Assays Average 2.75 Singida, Tanzania [press release], Available at

www.kilimanjarominingcompany.com/investors/news/2009/8-3.pdf, [6 December,

2011]

10. Lake Victoria Mining Company Inc., 2009, Lake Victoria Mining Company: 15 Additional

Gold Tailing Piles Average 3.03 Grams per Metric Ton [press release], Available at

www.infomine.com/index/pr/Pa790864.PDF, [6 December, 2011]

11. Mitchell, C. J. 1997, A Review of Gold Particle Size and Recovery Method.[pdf], Technical

Report WC/97/14, British Geological Survey

12. RGL, 2010, High Grade Copper-Gold Analysis Results for Burraga Tailing [press release],

Available at

http://www.republicgold.com.au/documents/announcements/522960_92846_High%20

Grade%20Copper-Gold%20Analysis%20Results%20Burraga%20Tailings.pdf, [10

December,2011]

[55]

October 2012

APPENDIX 1

Sample point location maps for Tailings Stockpiles

[56]

October 2012

Figure 14: Sample point layout for tailing BTA

Figure 15:Sample point layout for tailing BTB

[57]

October 2012

Figure 16: Sample point layout for tailing BTC

[58]

October 2012

Figure 17: Sample point layout for tailing BTD

Figure 18: Sample point layout for tailing BTE

[59]

October 2012

Figure 19: Sample point layout for tailing BTF

Figure 20: Sample point layout for tailing BTG

[60]

October 2012

Figure 21: Sample point layout for tailing BTH

[61]

October 2012

Figure 22: Sample point layout for tailing BTI

Figure 23: Sample point layout for tailing BTN

[62]

October 2012

APPENDIX II

Table showing sample point data

[63]

October 2012

Table 7: Summary of Sample Data

Tailing Tailing

Field ID

Sample

Tag No.

Tailing Coordinate

(PSAD 56) Sample Description From To

Hole

Depth

Assay

Results

Pile 1 BTA g/ton

1.1 952696 21 N 0302671 0602883 Orange soil + gravel 0 1 0

1.2 952697 " Greyish clayey material - Slush 1 2 2 0

2.1 952698 " Orange soil + grey clayey slush 0 1 0


3.1 952700 " Orange soil + grey clayey slush 0 1 1 0

4.1 952701 " Thin orange soil layer followed

by greyish slush

0 1 1 0


6.1 952703 " Orange soil followed by greyish

slush

0 1 1 0

7.1 952704 " Orange soil + grey clayey slush 0 1 0

7.2 952705 " Clayey slightly slushy material 1 2 2 0

NILL 952706 NILL Sample Does not Exist NILL

8.1 952707 " orange sandy soil 0 1 0

8.2 952708 " Greyish clayey material with

some sand

1 2 2 3

9 952709 " orange sandy soil 0 1 1 4

10 952710 " orange soil (thin) 0.2m 0 1 1 0

11 952711 " orange soil (thin) 0.2m 0 1 1 0

12.1 952712 " orange soil (thin) slush below 0 1 0

12.2 952713 " With sandy material + Gravel 1 2 2 0

Pile 2 BTB

1.1 952714 21 N 0302646 0602883 thin layer orange soil + slush 0

2.1 952715 " orange soil + gravel + greyish

sandy soil

0

[64]

October 2012

2.2 952716 " grey sandy soil followed by grey

clayey material

0

3.1 952717 " orange to soil (0.5m) followed

by grey-white sandy soil + gravel

0

4.1 952718 " orange top soil (1ft) followed by

mud soil composite (0.5ft)

0 0.5 0

4.2 952719 " slushy material 0.5 1 1 0

5.1 952720 " orange soil 0 0.3 0

5.2 952721 " soft clayey material + gravel 0.3 0.6 0.6 0

6.1 952722 " orange soil + gravel 0 0.3 0

6.2 952733 " grey sandy material + gravel (3 ft

thick)

0.3 1.22 1.22 0

7.1 952724 " orange soil 0 0.6 0

7.2 952725 " white sandy soil + gravel 0.6 1.22 0

7.3 952726 " creamy soft clay continue from

above

1.12 1.52 1.52 0


8.2 952728 " grey white sand with slush

below

0.6 1 1 0

9 952729 " orange soil up, slush below 0 0.45 0.45 0

10 952730 " white sandy soil + gravel (0-2ft),

slushy sandy material + gravel(2-

3ft)

0 1 1 0


11.2 952732 " white sand + gravel, slush below 0.6 1.22 1.22 0


12.2 952734 " white sandy soil with slush

below

0.6 1.22 1.22 0

13.1 952735 " white sandy soil(0-2 ft), sandy

clay + gravel, slush below

0.6 1.22 1.22 0

[65]

October 2012

Pile 3 BTC

1 952736 21 N 0302722 0602871 soft clayey material (approx.

1ft), slushy material (1-1.5ft)

0 0.76 0.76 0

2 952737 " clayey slushy material 0

3 952738 " soft clay - tailing slush 0 0.5 0.5 0

4 952739 " grey-white sand + gravel 0 1 1 0

5.1 952740 " grey-white sand + gravel 0 0.6 0

5.2 952741 " soft clayey material - partially

slush

0.6 1.22 1.22 0

6 952742 " Sandy material with slush below 0 0.6 0.6 0

7 952743 " grey white sandy soil 0 0.45 0.45 0

Pile 4 BTD

1.1 952744 21 N 0302733 060818 orange soil 0 0.6 0

1.2 952745 " grey sandy clay material 0.6 0.91 0

1.3 952746 " wet slushy clay material 0.91 1.52 1.52 4

2.1 952747 " orange top soil followed by grey-

white sand silt material

0 0.6096 2.67

2.2 952748 " wet greenish clayey material 0.6096 1.2192 0

2.3 952749 " wet grey-white sand silt material

(possibly top soil material from

underlying surface)

1.2192 1.8288 1.82 0

3.1 952750 " orange top soil followed by grey-

white sand silt material

0 1 1 0

4.1 952751 " Orange + grey white soil, bottom

half wet

0 1 1 0

5.1 952752 " Thin orange top soil followed by

grey white sandy soil

0 0.6096 0

5.2 952753 " wet grey-white sandy soil, clay

below

0.6096 1 1 0

6.1 952754 " Grey-white sandy soil 0 1 3

[66]

October 2012

6.2 952755 " clayey material + gravel and

sand (wet)

1 2 2 4

7.1 952756 " Orange soil 0 0.6096 0

7.2 952757 " grey white sandy clay + gravel 0.6096 1 1 0

8.1 952758 " grey white sandy clay (wet) 0 0.5 0.5 0

9.1 952759 " grey white sandy soil 0 0.6096 0

9.2 952760 " grey white sandy clay (wet) 0.6096 1.524 1.524 0

10.1 952761 " orange, grey white soil 0 1 0

10.2 952762 " Orange grey white sandy clay +

gravel (wet)

1 2 2 0

11.1 952763 " Orange grey white sandy soil 0 1 1 0

12.1 952764 " Grey white sandy soil + gravel 0 0.6096 0

12.2 952765 " Grey white sand clay material 0.6096 1.524 1.524 0

Pile 5 BTE

1.1 952766 21 N 0302702 0602818 grey white sandy soil + gravel 0 1 0

1.2 952767 " grey white sandy soil + gravel 1 2 2 0

2.1 952768 " grey white sandy soil 0 1 0

2.2 952769 " grey white sandy soil 1 2 2 0

3.1 952770 " grey white sandy soil 0 1 0


4.1 952772 " grey white sandy soil + gravel 0 1 0


5.1 952774 " grey white sandy soil + gravel

with clay below

0 1 0

5.2 952775 " grey white sandy soil 1 2 2 0

6.1 952776 " grey white sandy soil + gravel 0 0.6096 0


(wet)

0.6096 2 2 0

7.1 952778 " Grey white sandy material +

gravel

0 1 0

[67]

October 2012

7.2 952779 " grey white sandy material with

some clay

1 2 2 0

8.1 952780 " orange colored soil + gravel 0 0.3048 0


(wet )

0.3048 2 2 0

Pile 6 BTF

1.1 952782 21 N 0302655 0602840 orange and white sandy soil 0 1 0

1.2 952783 " orange and white sandy soil +

gravel

1 2 0

1.3 952784 " greenish colored sandy clay +

gravel

2 3 3 0

2.1 952785 " Greenish sandy clay, located 4 ft

below hole one. See diagram

0 0.6096 0.6096 0

3.1 952786 " grey white sandy material 0 1 0

3.2 952787 " grey white sandy material (wet) 1 2 2 0

4.1 952788 " grey white sand material +

gravel

0 1 1 0

5 952789 " grey white sand + gravel 0 0.6096 0.6096 3.67

6.1 952790 " grey white sand + gravel 0 1 0

4.2 952791 " grey white sandy material +

gravel

1 2 2 0

6.2 952792 " brown sandy material 1 2 2 43.67

7.1 952793 " orange and grey white sand +

gravel

0 0.6096 0

7.2 952794 " greenish sandy clay material 0.6096 1 1 0

8.1 952795 " orange and grey white sand +

gravel

0 1 0

8.2 952796 " greenish sandy clay (clay layer

below)

1 2 2 0

9.1 952797 " orange white sand material + 0 1 0

[68]

October 2012

gravel

9.2 952798 " greenish sand clay material 1 2 0

9.3 952799 " greenish sandy clay material 2 2.7 2.7 0

10.1 952800 " orange white sand 0 1 0

10.2 952801 " greenish sandy clay material 1 2 2 0

11.1 952802 " sandy clay material (clay layer

below)

0 1 1 0

12.1 952803 " grey white top soil followed by

sandy clay

0 1 1 0

13.1 952804 " grey white sandy material with

sandy clay below

0 1 1 0

14.1 952805 " grey white sandy soil with clay

material below

0 1 1 0

15.1 952806 " grey white sand with clay below,

followed by slush

0 1.5 1.5 0

16.1 952807 " grey white sand material +

gravel (wet)

0 1 1 0

Pile 7 BTG

1.1 952808 21 N 0302619 0602807 top grey-white sand, clay

material below

0 1 0

1.2 952809 " clayey material 1 1.2192 1.2192 0

2.1 952810 " grey white sand followed by clay

slush below

0 1 1 0


3.2 952812 " green-gray clay material, slush

below

1 1.2192 1.2192 0

4.1 952813 " orange colored soil + gravel 0 1 0

4.2 952814 " greenish colored sandy clay

material

1 2 2 0

5.1 952815 " orange sandy soil 0 1 2.83

[69]

October 2012

5.2 952816 " grey sandy clay material 1 2 2

5.3 952817 " grey clayey material (wet) 2 3 3 0

6.1 952818 " orange sandy top soil with sandy

clay below

0 1 1 1.3

7.1 952819 " orange sandy clay material , logs

below

0 0.6096 0.6096 0

8.1 952820 " orange sandy soil material 0 1 0

8.2 952821 " greenish sandy clay material 1 2 0

8.3 952822 " grey sandy clay material 2 3 3 0

9.1 952823 " orange sandy top soil with clay

material below

0 1 0

9.2 952824 " orange grey sand clay material 1 2 0

9.3 952825 " orange grey material + gravel 2 2.7 2.7 0

10.1 952826 " orange sandy material 0 1 0

10.2 952827 " grey-green sand clay material 1 1.7 1.7 0

11.1 952828 " orange sandy material (wet) 0 1 1 0

12.1 952829 " orange grey sand clay material 0 1 1 0

Pile 8 BTH

1.1 952830 21 N 0302692 0602719 orange soil with grey white clay

below

0 1 1 0

2.1 952831 " orange soil with clay below 0 0.762 0.762 0

3.1 952832 " orange soil with sandy clay

below

0 1 0

3.2 952833 " clayey material, grey-white sand

clay at top

1 1.5 1.5 0

4.1 952834 " orange soil with grey-white clay

material below

0 1 0

4.2 952835 " orange greyish clay material 1 2 2 0

5.1 952836 " orange soil + gravel 0 1 0

5.2 952837 " orange grey-white sandy soil + 1 1.7 1.7 0

[70]

October 2012

gravel


6.2 952839 " grey-white sandy material 0.6096 1 0

6.3 952840 " grey-white sandy clayey material

(wet)

1 2 0

6.4 952841 " clayey material + sand material 2 2.1336 2.1336 0

7.1 952842 " orange soil with grey-white soil

below

0 1 0

7.2 952843 " grey white sandy material

+gravel

1 2 2 0

8.1 952844 " orange soil 0 1 0

8.2 952845 " grey-white soil material + gravel 1 2 2 0

9.1 952846 " orange soil 0 0.762 0

9.2 952847 " grey-white sandy material with

clay material

0.762 2 2 5.3

10.1 952848 " orange top soil followed by grey

white soil

0 1 2

10.2 952849 " grey greenish soil (clayey

material)

1 2.5 2.5 4

11.1 952850 " orange top soil with grey-white

material below

0 1 0

11.2 952851 " greenish clayey material 1 1.6 1.6 0

12.1 952852 " orange top soil followed by

sandy clay

0 1 1 0

13.1 952853 " orange sandy soil 0 1 0

13.2 952854 " orange sand clay material 1 1.7 1.7 3.4

14.1 952855 " orange sandy soil followed by

grey-white sandy material

0 1.5 0

14.2 952856 " orange colored sand clay

material

1.5 2 2 0

[71]

October 2012

15.1 952857 " Orange sandy soil 0 1 0

15.2 952858 " grey white sand with clay

material below

1 2 2 0


16.2 952860 " orange clayey material 1 1.5 1.5 0


17.2 952862 " greenish clayey material 1 2 2 0

Pile 9 BTI

1.1 952863 21 N 302665 602701 orange sandy soil 0 1 0

1.2 952864 " grey white sandy soil with a

clayey mix below

1 2 0

1.3 952865 " sandy clay material + gravel 2 3.5 3.5 0

2.1 952866 " orange sandy soil, white sandy

soil below

0 1 0

2.2 952867 " orange clayey sand material +

gravel

1 1.5 1.5 0

3.1 952868 " grey with sand + gravel 0 1 0

3.2 952869 " grey white sand continue 1 2 2 0

4.1 952870 " orange sandy soil with grey

white sand below

0 1 0

4.2 952871 " grey white sandy material 1 2 2 0

5.1 952872 " brown sand 0 1 2.5

5.2 952873 " grey sand 1 2 2 3

6.1 952874 " 0 1 0

6.2 952875 " 1 1.5 1.5 0

7.1 952876 " greyish brown sand 0 1 0

7.2 952877 " brown sand 1 2 2 0

8.1 952878 " brown sand 0 1 0

8.2 952879 " greyish brown sand 1 2 2 0

9.1 952880 " brown sand 0 1 0

[72]

October 2012

9.2 952881 " 1 2 2 2

10.1 952882 " 0 1 0


11.1 952884 " greyish brown sand 0 1 0


12.1 952886 " grey sand 0 1.5 1.5 0

13.1 952887 " grey sand 0 0.75 0.75 0

14.1 952888 " greyish brown sand 0 1.5 1.5 0

15.1 952889 " 0 1 1 0

16.1 952890 " 0 0.75 0.75 0

Pile

10

BTN

1 952911 21 N 302667 602739 Brown sand + quartz gravel 0 1 1 6

2 952912 " Brown sand + quartz gravel 0 1.2 1.2 2.5

3 952913 " Brown sand + quartz gravel 0 1.5 1.5 0



6 952938 " Brown sand + quartz gravel 0 1 0

6.2 952939 " Brown sand + quartz gravel 1 2 2 0

7 952940 " Brown sand 0 1 0


8 952942 " brown sand 0 1.5 1.5 1

9 952943 " Brown sand + quartz gravel 0 1 1 0

10 952944 " Brown sand + quartz gravel 0 1 0



[73]

October 2012

APPENDIX III

Gold Distribution Models of Tailings

Note that the outlines of the models do not necessarily represent the actual shape of the

tailings bodies.

[74]

October 2012

Figure 24: Gold Distribution Model for BTA

[75]

October 2012

Figure 25: Gold Distribution Model for BTD

[76]

October 2012

Figure 26: Gold Distribution Model for BTF

[77]

October 2012

Figure 27: Gold Distribution Model for BTG

[78]

October 2012

Figure 28: Gold Distribution Model for BTH

[79]

October 2012

Figure 29: Gold Distribution Model for BTI

[80]

October 2012

Figure 30: Gold Distribution Model for BTN

[81]

October 2012

APPENDIX IV

Operation profitability Charts

[82]

October 2012

Figure 31: Profitability Chart for the operation at 40% efficiency for gold at $1200 USD/Oz.

$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80

40% RECOVERY AT 1200 USD/Oz

Profitable Production

Accum. Cost

[83]

October 2012


$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80

US

Do

lla

rs

Days



Accum. Cost

[84]

October 2012


$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

$80,000.00

$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

$80,000.00

5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80

Days

US

Do

lla

rs

US

Do

lla

rs60% RECOVERY AT 1200 USD/Oz


Accum. Cost

[85]

October 2012


$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

$80,000.00

$90,000.00

$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

$80,000.00

$90,000.00

5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80

US

Do

lla

rs

US

Do

lla

rs

Days



Accum. Cost

[86]

October 2012


$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

$80,000.00

$90,000.00

$100,000.00

$0.00

$10,000.00

$20,000.00

$30,000.00

$40,000.00

$50,000.00

$60,000.00

$70,000.00

$80,000.00

$90,000.00

$100,000.00

5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80

US

Do

lla

rs

US

Do

lla

rs

Days



Accum. Cost

[87]

October 2012


$0.00

$20,000.00

$40,000.00

$60,000.00

$80,000.00

$100,000.00

$120,000.00

$0.00

$20,000.00

$40,000.00

$60,000.00

$80,000.00

$100,000.00

$120,000.00

5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80

US Dollars

US

Do

lla

rs

Days



Accum. Cost

[88]

October 2012

APPENDIX V

Definition of Terms

[89]

October 2012

Inferred, Indicated and Measured Mineral Resource according to NI43-101

“In this Instrument, the terms "mineral resource", "inferred mineral resource", "indicated

mineral resource" and "measured mineral resource" have the meanings ascribed to those

terms by the Canadian Institute of Mining, Metallurgy and Petroleum, as the CIM Definition

Standards on Mineral Resources and Mineral Reserves adopted by CIM Council, as those

definitions may be amended.”

The terms Measured, Indicated and Inferred are defined by CIM (2005) as follows:

“A Mineral Resource is a concentration or occurrence of diamonds, natural solid inorganic

material, or natural solid fossilized organic material including base and precious metals, coal

and industrial minerals in or on the Earth’s crust in such form and quantity and of such a grade

or quality that it has reasonable prospects for economic extraction. The location, quantity,

grade, geological characteristics and continuity of a Mineral Resource are known, estimated or

interpreted from specific geological evidence and knowledge.”

“The term Mineral Resource covers mineralization and natural material of intrinsic economic

interest which has been identified and estimated through exploration and sampling and within

which Mineral Reserves may subsequently be defined by the consideration and application of

technical, economic, legal, environmental, socio-economic and governmental factors. The

phrase ‘reasonable prospects for economic extraction’ implies a judgment by the Qualified

Person in respect of the technical and economic factors likely to influence the prospect of

economic extraction. A Mineral Resource is an inventory of mineralization that under

realistically assumed and justifiable technical and economic conditions might become

economically extractable. These assumptions must be presented explicitly in both public and

technical reports.”

Inferred Mineral Resource

“An ‘Inferred Mineral Resource’ is that part of a Mineral Resource for which quantity and grade

or quality can be estimated on the basis of geological evidence and limited sampling and

reasonably assumed, but not verified, geological and grade continuity. The estimate is based on

[90]

October 2012

limited information and sampling gathered through appropriate techniques from locations such

as outcrops, trenches, workings and drill holes.”

“Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be

assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or

Measured Mineral Resource as a result of continued exploration. Confidence in the estimate is

insufficient to allow the meaningful application of technical and economic parameters or to

enable an evaluation of economic viability worthy of public disclosure. Inferred Mineral

Resources must be excluded from estimates forming the basis of feasibility or other economic

studies.”

Indicated Mineral Resource

“An ‘Indicated Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or

quality, densities, shape and physical characteristics can be estimated with a level of confidence

sufficient to allow the appropriate application of technical and economic parameters, to

support mine planning and evaluation of the economic viability of the deposit. The estimate is

based on detailed and reliable exploration and testing information gathered through

appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes

that are spaced closely enough for geological and grade continuity to be reasonably assumed.”

“Mineralization may be classified as an Indicated Mineral Resource by the Qualified Person

when the nature, quality, quantity and distribution of data are such as to allow confident

interpretation of the geological framework and to reasonably assume the continuity of

mineralization. The Qualified Person must recognize the importance of the Indicated Mineral

Resource category to the advancement of the feasibility of the project. An Indicated Mineral

Resource estimate is of sufficient quality to support a Preliminary Feasibility Study which can

serve as the basis for major development decisions.”

[91]

October 2012

Measured Mineral Resource

“A ‘Measured Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or

quality, densities, shape, and physical characteristics are so well established that they can be

estimated with confidence sufficient to allow the appropriate application of technical and

economic parameters, to support production planning and evaluation of the economic viability

of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing

information gathered through appropriate techniques from locations such as outcrops,

trenches, pits, workings and drill holes that are spaced closely enough to confirm both

geological and grade continuity.”

“Mineralization or other natural material of economic interest may be classified as a Measured

Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of

data are such that the tonnage and grade of the mineralization can be estimated to within close

limits and that variation from the estimate would not significantly affect potential economic

viability. This category requires a high level of confidence in, and understanding of, the geology

and controls of the mineral deposit.”