Scavenging Flotation Tailings using a Continuous Centrifugal Gravity Concentrator

by

Hassan Ghaffari

B.A.Sc. & M.A.Sc. Department of Mining Engineering, Technical Faculty Tehran University, Iran, 1990

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Summary

A study was conducted to evaluate the Knelson Continuous Variable Discharge (CVD) concentrator as a scavenger for coarse middling particles from flotation tailings. The goal was to recover a product of suitable grade for recycling to the grinding circuit to improve liberation and aid subsequent recovery in flotation. Such a hybrid flotation- gravity circuit would result in improved metal recoveries, product grades and potentially lead to lower grinding costs.

Froth flotation, a physico-chemical process, is the most commonly used process for treating base metal sulphides. Since separation is achieved on the basis of the surface hydrophobicity, mixed-phased particles or middlings are not efficiently treated by froth flotation. The coarser fractions in flotation feeds contain heavy and valuable liberated and non liberated minerals that cannot be floated efficiently. There is a top limit for particle size floatability which varies for different ores. The specific gravity of these particles plays an important role which causes decreasing the flotation performance.

This study is based on the premise that the middling particles in tailings can be recovered efficiently by size enhanced density separation. The ability of gravity separators to treat fine particles has been limited by the lack of particle inertia relative to the surface drag forces. Particle inertia can be enhanced by the application of a centrifugal field. The Knelson CVD is a relatively new technology as a continuous centrifugal concentrator that recovers particles based primarily on density but also on size.

Tailing samples from two mines were subjected to characterization, batch gravity and pilot scale CVD testing. Table 1 shows the specifications of the analyzed samples. Ill

Table 1. Specification of the Analyzed Samples

Grade Mine Sample Au (ppm) S% Ni% Eskay Creek (Barrick) Low Grade Tails (LGT) l.l 1.0 N/A High Grade Tails (HGT) 1.9 1.2 N/A Birchtree (INCO Thompson) Rougher-Scavenger Tails (AS-4) <1 3.0 0.2 Cleaner-Scavenger Tails (SC-4) <1 4.2 0.5

The sample characterization of the both Eskay Creek and INCO Thompson samples indicated that there is potential to upgrade gold and nickel in the coarsest fractions. A preliminary assessment of gold and nickel recovery was obtained using a batch Knelson concentrator and the results were similar to those predicted from the characterization results.

Based on the characterization and batch gravity tests, there appeared to be a greater opportunity to recover a recyclable product from the LGT and the SC-4, samples than the HGT and AS-4 samples. They were therefore selected for pilot scale tests using the CVD6 separator. The objective was to produce a concentrate with a low mass yield and high gold and nickel upgrade ratios. Products from these tests were subjected to size/assay analysis.

The results for Eskay Creek sample show that the gold distributions are significantly higher than sulphur and arsenic distributions, which indicates that free gold or gold associated with poorly floated minerals is being recovered. The best result was obtained from Test 3 in which 28% of the gold was recovered in a concentrate grading 12.5 ppm Au representing 3.3% of the feed mass. This represents a gold upgrade ratio of 8.7.

The results for INCO Thompson show that the gold distributions are significantly higher than nickel and sulphur distributions, which indicates that free gold is being recovered. The best result was obtained from Test 3 in which 12.9% of the nickel was recovered in a concentrate grading 0.85% Ni representing 6.7% of the feed mass. It is iv worth noting that the Mg grade is decreased dramatically in the concentrate products. The Mg distribution in concentrate was just 1.7%. This is a good indication that the centrifugal concentrator can also remove the magnesium bearing minerals such as talc. In the case of gold, remarkable results were produced and the metallurgical balances showed gold recoveries ranging from 12.6 to 67.3% and gold grades from 1.4 to 10.0 ppm.

Size partition curves were obtained for all tests, which demonstrated that the CVD operates as a size classifier as well as a density separator (size enhanced density separation). For the Eskay sample (Tests 3 and 6) and the INCO sample (Test 7), the cut size was about 300 microns.

In conclusion, the pilot scale testing indicated that the CVD was effective at recovering gold, gold bearing sulfides and nickel bearing sulfides from the coarse particles of the flotation tailings and also at de-sliming. The nickel and gold recovered were primarily in the coarse size fractions, which would likely contain middling particles (sulfides + silicates). The results showed that the CVD is capable of rejecting the Mg bearing minerals as well. These results support the application of continuous centrifugal concentrators into hybrid flotation-gravity circuits that could lead to improved metallurgical performance.

Plant trials testing are recommended to confirm the results and potential benefits. V

TABLE OF CONTENTS

Summary ii

Table of Contents v

List of Figures vii

List of Tables x

Acknowledgements xiii

CHAPTER 1 Introduction 1

1.1. Background 1 1.2. Methodology 2

CHAPTER 2 Literature Review 4 2.1. Introduction 4 2.2. Relationship between Mineral Recovery and Particle Size 4 2.3. Enhanced Gravity Separation Process Technologies 7 2.4. Continuous Centrifugal Gravity Concentrators 8 2.4.1. Types 8 2.4.2. Applications 14 2.5. Conclusion 22

CHAPTER 3 Experimental Program 23

3.1. Analyzed Samples 23 3.1.1 Eskay Creek 23 3.1.2. INCO Thompson 24 3.2. Sample Characterization 25 3.2.1. Sample Preparation 25 3.2.2. Size/Assay Analysis 26 3.2.3. Mineralogical characterization 26 3.2.4. Density Fractionation 26 3.3. Batch Tests 28 3.4. Pilot Scale Tests 29 3.4.1. Products Characterization 3 0 3.4.2. Concentrate Density Fractionation Tests 31 3.4.3. Flotation Test and Proposed Flowsheet 31

CHAPTER 4 Results and Discussion 32

4.1. Eskay Creek Samples 32 4.1.1. Introduction 32 vi

4.1.2. Results and Discussion 33 4.1.3. Conclusion and Recommendation 47 4.2. INCO Thompson sample 49 4.2.1. Introduction 49 4.2.2. Results and Discussion 50 4.2.3. Conclusion and Recommendation 70

CHAPTER 5 Conclusions and Recommendations 73

5.1. Conclusions 73

5.2. Recommendations 75

REFERENCES 76

APPENDICES 81 Appendix I - Eskay Creek 81 Appendix IA Sample Characterization 82 Appendix IB Batch Tests 86 Appendix IC Pilot Scale Test Conditions and Results 87 Appendix ID Mineralogical Analysis 101 Appendix IE Flotation Test 104

Appendix II - INCO Thompson 105 Appendix IIA Sample Characterization 106 Appendix IIB Batch Tests 117 Appendix IIC Pilot Scale Test Conditions and Results 118 Appendix IID Density Fractionation Test Results 132 Appendix HE Flotation Test 141 List of Figures

Figure2.1 Dependence of flotation efficiency on particle size for several ores 6

Figure2.2 Typical grinding throughput, product size and flotation Recovery relationship for a base metal concentrator 7

Figure2.3 Continuous Centrifugal Gravity Concentrators 9

Figure 3.1 Eskay Creek Flotation Circuit Sample Point 23

Figure 3.2 Birchtree Flotation Circuit Sample Points 24

Figure 3.3 Sampling procedure of wet tailing samples 25

Figure 3.4 Batch Knelson test procedure 28

Figure 4.1 Au Grade Distributions, LGT & HGT 34

Figure 4.2 S Grade Distributions, LGT & HGT 34

Figure 4.3 Au Grade Distributions in Concentrate and Tailing of Test 3, LGT 38

Figure 4.4 S Grade Distributions in Concentrate and Tailing of Test 3, LGT 39

Figure 4.5 Mass Yield and Au Recovery versus Pinch

Valve Closed Time, LGT 40

Figure 4.6 Au Grade/Recovery versus Mass Yield, LGT 41

Figure 4.7 Au Upgrade Ratio/Recovery versus Mass Yield, LGT 41

Figure 4.8 Size Separation Partition Curves, LGT 42

Figure 4.9 Au Recovery versus Size, LGT 43

Figure 4.10 S Recovery versus Size, LGT 44

Figure 4.11 Au Recovery versus S Recovery, LGT 44

Figure 4.12 Proposed Flowsheet, Eskay Creek 46

Figure 4.13 Ni Grade for each Size Fraction, SC-4 & AS-4 51 Vlll

Figure 4.14 Mg Grade for each Size Fraction, SC-4 & AS-4 52

Figure 4.15 S Grade for each Size Fraction, SC-4 & AS-4 52

Figure 4.16 Ni Grade for each Size Fraction of Concentrate & Tailing of Test 3, SC-4 56

Figure 4.17 Mg Grade for each Size Fraction of Concentrate & Tailing of Test 3, SC-4 57

Figure 4.18 S Grade for each Size Fraction of Concentrate & Tailing of Test 3, SC-4 57

Figure 4.19 Au Grade for each Size Fraction of Concentrate & Tailing of Test 3, SC-4 58

Figure 4.20 Mass Pull, Ni Grade/Recovery versus Pinch Valve Closed Time, SC-4 59

Figure 4.21 Ni Upgrade and Mg Downgrade Ratios versus

Mass Pull, SC-4 60

Figure 4.22 Ni Grade/Recovery versus Cone. Mass Pull 60

Figure 4.23 Size Separation Partition Curves, SC-4 62

Figure 4.24 Partition Coefficient Curve for +50# Size Fraction 63

Figure 4.25 Partition Coefficient Curve for -50+100# Size Fraction 64

Figure 4.26 Ni Recovery for +50# Size Fraction 64

Figure 4.27 Ni Recovery for -50+100# Size Fraction 65

Figure 4.28 Mg Recovery for -100+140# Size Fraction 65

Figure 4.29 Mg Recovery for-140+200# Size Fraction 66

Figure 4.30 S Recovery for +50# Size Fraction 67

Figure 4.31 Fe Recovery for +50# Size Fraction 67

Figure 4.32 Proposed Flowsheet, Inco Thompson 69

Figure A. 1.1 Particle Size Distribution, LGT 82 Figure A. 1.2 Particle Size Distribution, HGT 83

Figure C.6.1 Size Separation Partition Curves, LGT 101

Figure C.6.2 Au Recovery versus Size, LGT 101

Figure C.6.3 S Recovery versus Size, LGT 102

Figure C.6.4 Au Recovery versus S Recovery, LGT 102

Figure C.6.5 Size Separation Partition Curves, SC-4 138

Figure C.6.6 Partition Coefficient Curves for each Size Fraction (Mass Yield versus Density) 138

Figure C.6.7 Ni Recovery for each Size Fraction (Ni Recovery versus Density) 139

Figure C.6.8 Mg Recovery for each Size Fraction (Mg Recovery versus Density) 139

Figure C.6.9 S Recovery for each Size Fraction (S Recovery versus Density) 140

Figure C.6.10 Fe Recovery for each Size Fraction (Fe Recovery versus Density) 140 List of Tables

Table 1 Specification of the Analyzed Samples iii

Table 2.1 Continuous Centrifugal Gravity Concentrators Technologies and Applications 14

Table 2.2 Falcon Model C Operating Applications 15

Table 2.3 The Kelsey Jig Operating Applications 17

Table 2.4 Tested applications for the Altair Centrifugal Jig 19

Table 2.5 The Knelson CVD Tested and Operating Applications 21

Table 3.1 Tests Operating Conditions, LGT 29

Table 3.2 Tests Operating Conditions, SC-4 30

Table 4.1 Ore Zones Assays (1998), Eskay Creek 32

Table 4.2 Tailing Samples and Ore Types Specifications

(Eskay Creek Plant, 2003) 33

Table 4.3 Samples Characterization Results, LGT & HGT 35

Table 4.4 Degree of Liberation Results, LGT 36

Table 4.5 Batch Test Results, LGT & HGT 36

Table 4.6 Pilot Scale Test Results, LGT 37

Table 4.7 Flotation Test Results, LGT 45

Table 4.8 Birchtree Ore Mineralogy 49

Table 4.9 Birchtree Ore, Concentrate and Tailing Grades 49

Table 4.10 Samples Characterization Results, SC-4 & AS-4 53

Table 4.11 Batch Test Results, SC-4 & AS-4 54

Table 4.12 Pilot Scale Test Results, SC-4 55

Table 4.13 Flotation Test Results, SC-4 68 XI

Table A. 1.1 LGT Size Analysis 82

Table A. 1.2 HGT Size Analysis 83

Table A. 1.3 Specification of Low Grade Tails (LGT) and Mill Performance (Eskay Creek Plant Results, 2003) 84

Table A. 1.4 Specification of High Grade Tails (HGT) and

Mill Performance (Eskay Creek Plant Results, 2003) 84

Table A.2.1 Size/assay Analysis, LGT 85

Table A.2.2 Size/assay Analysis, HGT 85

Tables C.3.1 to C.3.6 CVD Test Products, Size Analysis and PN, LGT 89

Tables C.4.1 to C.4.6 CVD Test Products, Concentrate Size/ Assay Analysis, LGT 92 Tables C.5.1 to C.5.6 CVD Test Products, Tailings Size/ Assay Analysis, LGT .95

Tables C.6.1 to C.6.6 CVD Test Products, Size/ Assay

Analysis and Mass Yield, LGT 98

Table A.3.1 Magnetic Particles Distribution, SC-4 109

Table A.3.2 Magnetic Particles Distribution, AS-4 110

Table A.3.3 Magnetic Separation Products Size/Assay Results, SC-4 111

Table A.3.3 Magnetic Separation Products Size/Assay Results, AS-4 112

Table A.4.1 Magstream Products, SC-4 Head Sample 113

Table A.4.2 Magstream Products, AS-4 Head Sample 115

Table B.l Knelson Batch Test Results SC-4 117

Table B.2 Knelson Batch Test Results AS-4 117

Table C. 1 CVD Test Conditions SC-4 118

Table C.2 Metallurgical Balances SC-4 119

Tables C.3.1 to C.3.8 CVD Test Products, Size Analysis and PN, SC-4 120 Tables C.4.1 to C.4.7 CVD Test Products, Concentrate Size/ Assay Analysis, SC-4

Tables C.5.1 to C.5.7 CVD Test Products, Tailings Size/ Assay Analysis, SC-4

Table D.l Magstream Products, CVD Test 7 Concentrate, SC-4

Tables D.2.1 to D.2.8 Partition Coefficient Numbers and Metal Recoveries, SC-4 xiii

Acknowledgements

The work presented here was funded by: - Natural Sciences and Engineering Research Council of Canada (IPS Program) - Knelson Concentrators The sampling and shipment was done by: - INCO Manitoba Division, Birchtree Mine - Barrick Gold Corp., Eskay Creek Mine The work was directed and supervised by: Dr. Bern Klein The advice and support of all is greatly acknowledged. CHAPTER 1 Introduction

1.1. Background

Froth flotation is the main industrial process that is used to recover metal minerals from deposits. Metal recoveries from froth flotation range typically from 70% to 95%. Metal

losses are primarily in either the coarsest or finest particles. The coarsest fractions

contain middling particles that have insufficient metal minerals content to float or are too

coarse to be recovered efficiently by flotation.

Continuous centrifugal concentrators are a relatively new technology that evolved

from batch centrifugal concentrators that are used primarily for recovering gold particles.

The main continuous centrifugal gravity concentrators that are commercially available

include Knelson Continuous Variable Discharge (CVD), Falcon Model C, Kelsey Jig,

Altair Centrifugal Jig and Mozley Multi-Gravity Separator (MGS).

The CVD has been shown to recover particles by density, particularly at coarse

sizes. Therefore, a study was conducted to assess its ability to recover coarse middlings

into a product suitable for recycling to grinding. It is proposed that this hybrid flotation-

gravity concentration circuit could lead to improved metal recoveries, higher concentrate

grades and lower grinding requirements.

The specific objectives of this study are as follows:

• To evaluate the CVD concentrator as a scavenger for coarse middling

particles from flotation tailings

• To identify operating conditions with respect to the objective of achieving

high metal recoveries in a small mass yield.

• To assess size/density classification in CVD 2

To achieve these objectives, samples of flotation tailings from the Barrick Eskay

Creek and the INCO Thompson beneficiation plants were subjected to batch and pilot scale tests.

1.2. Methodology

The Eskay Creek samples are referred to as High Grade Tails (HGT, 1.9 ppm Au) and

Low Grade Tails (LGT, 1.1 ppm Au). The Thompson samples are Scavenger-Cleaner

Tails (SC-4, 0.5% Ni) and Rougher-Scavenger Tails (AS-4, 0.2% Ni) from processing

Birchtree ore. The test program for each sample was directed into three stages:

1. Sample characterization

2. Batch gravity testing

3. Continuous pilot scale testing

In the first stage, the samples were characterized to identify metal losses from flotation and to assess potential recoveries associated with coarse middling particles. For stage two, testing was performed using a batch Knelson concentrator to obtain a preliminary assessment of metal recovery.

Based on the characterization and batch testing, continuous pilot testing was performed to assess metallurgical performance.

The thesis outline is as follows:

• Chapter 1 is an introduction to research, objectives and methodology

• Chapter 2 reviews the literature and describes continuous centrifugal

gravity concentrator technologies.

• Chapter 3 describes the experimental program including the

characterization procedures and test conditions.

• Chapter 4 presents and discusses the test results Chapter 5 comprises of conclusions and recommendations 4

CHAPTER 2 Literature Review

2.1. Introduction

Froth flotation is a physico-chemical process and is the main industrial process that is used to recover base metal minerals from ore. It has been shown that froth flotation is not effective at recovering coarse particles that are either middling particles that have insufficient metal mineral content or can not be carried to the froth due to shear forces between the bubbles and particles. Since separation is achieved on the basis of surface hydrophobocity, mixed phase particles or middlings are not efficiently treated by froth flotation (Honaker et al., 1995).

Middling particles contain grains of the desired mineral and grains of other minerals. If many middling particles report to the concentrate, the metal grade will decrease and, if these particles report to the tailing, the metal recovery will decrease. In order to achieve a reasonable mineral recovery, the grinding and flotation circuits should be examined simultaneously which requires detailed process mineralogy to characterize the ore and products.

2.2. Relationship between mineral recovery and particle size

The efficiency of froth flotation in mineral processing depends on the size of the mineral, gangue and middling particles. The dependence of flotation on particle size for several ores is shown in Figure 2.1 (Reinson data are from Goodman et al., 1977 and Broken Hill

South data are from Cameron et al., 1965). The plots indicate that the mineral recovery has a maximum value over a range of particle size with a decline at the coarse and fine ends. According to Figure 2.1, recoveries begin to decrease for particles coarser than about 100 um which varies depending on the specification of the operation.

Improvements within the flotation circuit itself have the potential to shift the entire recovery curve (Figure 2.1) to higher values and extend the recovery over a wider 5 size range. The other opportunity is to seek to optimize the grinding circuit to place more of the product in the size range of maximum recovery. This is not an easy task (Napier-

Munn et al., 1999).

Applying lower grinding throughput or a finer hydrocyclone cut-size can reduce the coarse material fractions in the grinding circuit but on the other hand it increases the undesirable fine material fractions. Multi-stage cycloning has been tested as one approach to improving the overall distribution of flotation feed. 6

Figure 2.1. Dependence of flotation efficiency on particle size for several ores (after Lynch et al., 1981)

Generally, the coarser the material in flotation feed the lower is mineral recovery in the concentrate. Mineral recovery is related to the proportion of the coarse material in the feed. Figure 2.2 shows the relationship between grinding throughput, product size, and flotation recovery. The grinding circuit regimes determine the flotation feed size; with increasing the grinding circuit throughput, the percentage of coarse fraction increases. As Figures 2.1 and 2.2 represent, with increasing the coarse fraction in 7 flotation feed, the metal recovery decreases significantly. To determine the best condition for throughput, grind and recovery is not simple.

% coarse fraction Grinding throughput (tph) In Dotation feed

Figure 2.2. Typical grinding throughput, product size and flotation recovery relationship for a base metal concentrator (Napier-Munn, 1999)

According to Figure 2.1, recoveries also begin to decrease for particles finer than

about 20 urn. In order to reduce the generation of these ultrafine particles and metal

losses, coarse grinding and flotation followed by size enhanced gravity separation may be

an alternative to common processes. Scavenging flotation tailings using a continuous

centrifugal gravity separator may recover the heavy coarser particles. The goal is to

recover a product of suitable grade for recycling to the grinding-flotation circuit to

improve liberation and aid subsequent recovery in flotation. If good recoveries can be

achieved using the CVD as a scavenger, there is opportunity to use a coarser grind to

reduce the power requirements and associated costs. This will also reduce the amount of

fine (-20 um) material that floats poorly.

2.3. Enhanced Gravity Separation Process Technologies

During the past decade, several enhanced gravity concentrators have evolved to

commercial scale as a result of the need to achieve efficient separation of finer particles.

The initial interest in enhanced gravity separators was for the recovery of fine gold,

which represents a low product flow rate application whereby the concentrated gold

stream represents 1 % or less of the total feed. As a result, semi-batch units are the most 8 common enhanced gravity separator used in gold applications, in which, the feed and overflow streams are continuous while the gold particles are collected against the wall of a spinning bowl and periodically flushed to a concentrate launder. However, due to the need to treat materials having high-density particle contents greater than 1%, fully continuous units have been developed (Mohanty et al., 1999).

Extending the particle size limits of devices such as spirals, improved metallurgical performance over non-sulphide mineral flotation and pre-concentration are incentives to develop the centrifugal gravity concentrators further. Several enhanced centrifugal gravity separators including the batch semi-continuous and continuous concentrators have been designed and commercialized during the past decade for treating the fine particles. The ability to remove concentrate continuously extends the application of centrifugal concentrators requiring larger mass pulls, such as for the recovery of gold bearing and metal sulphide minerals.

The main reasons for selecting continuous technology are as follows:

• When target metal or mineral is present in percentage rather than ppm

concentrations, the fluidized rings become loaded with heavy mineral or metal

very quickly.

• Batch technologies must shut down to remove concentrate

• They are ideal for a wide range of mass yield application

(Simpson, 2003)

2.4. Continuous Centrifugal Gravity Concentrators

2.4.1. Types

The main enhanced gravity concentrators that operate continuously and are commercially available include the Falcon Model C, the Mozley Multi-Gravity Concentrator (MGS), the Kelsey Jig, the Altair Centrifugal Jig, and the Knelson Continuous Variable

Discharge (Knelson CVD). Figure 2.3 shows photos of Falcon C and Knelson CVD. 9

Falcon Model C Knelson CVD

Figure 2.3. Continuous Centrifugal Gravity Concentrators

2.4.1.1. Falcon Model C

The Falcon Model C is a continuous centrifugal separator. The operation is similar to

that of the Model SB (Batch concentrator), except the bowl shape is tapered facilitating

upward flow along the wall. No feed interruptions are required and the high density

product is de-slimed and partially de-watered.

In this machine, feed enters through the top and travels down to the bottom of a

rotational bowl where centrifugal acceleration forces particles to the wall. The particles

travel up the bowl section where heavy particles displace light particles along the bowl

wall. At the top of the bowl is a concentrate collection ring with valves positioned radially at the back of the ring. The valve aperture size is controllable and remains open.

The tailing material forms the innermost layer on the bowl wall and overflows the bowl to a tailings launder (Falcon, 1999; Silva et al., 1999).

The maximum feed solid content is 45%; however unlike the other continuous concentrators that are described, no fluidization/hutch water is required. This machine has only two operating variables; bowl speed and valve aperture size. Advantages of the 10

Falcon Model C compared to other continuous centrifugal concentrators are its high capacity, its relative mechanical simplicity and its ability to process particles up to 1 mm.

2.4.1.2. Kelsey Jig

The Kelsey Centrifugal Jig is an enhanced gravity separation device, which has the ability to increase the apparent gravitational field acting on fine feed particles across a bed of ragging material. Utilizing the same basic parameters as a conventional jig, as well as the enhancement of the gravitational field (achieved by spinning the jigging mechanism), enables very efficient separation of both fine minerals and minerals with a relatively small specific gravity differential (Richards et al., 2004).

The Kelsey Jig is fed down a fixed central pipe and feed is distributed upward over the surface of a bed of ragging material, which is supported by a cylindrical shaped screen. The screen is spun with the rotor and pressurized water is introduced into a series of hutches behind the screen. This water is pulsed through the ragging particles facilitating stratification. The denser particles pass through the internal screen to concentrate hutch and then through spigots to a concentrate launder, while the lighter mineral particles are discharged over a ragging retention ring into a tailing launder. The three controllable variables in the Kelsey Jig are bowl speed, ragging size/density and pulsation stroke length (Geologies, 1999; Silva et al., 1998; Wyslouzil, 1990)

The Kelsey Jig combines centrifugal acceleration (up to 50 G) with jigging action.

Changes in the design have improved its mechanical reliability. The development of the

J1800, with a capacity to 65 tph, has extended the potential applications. Feed must be screened to remove oversize (+50 mesh). Feed slurry solid contents of up to 65% can be processed; however pulse water dilutes the slurry to less than 50% solids.

The models available include:

• J200 KCJ - laboratory test unit, with nominal capacity of 15-100 kg/h solids 11

• J1300 Mkll KCJ - smallest commercial unit, with nominal capacity of 2-30 t/h

solids

• J1800 KCJ - largest commercial unit, with nominal capacity of 5-60 t/h solids

(Roche Mining, 2004)

2.4.1.3. Altair Centrifugal Jig

The Altair Centrifugal Jig is an enhanced gravity separator that provides relatively efficient separation on +25 um size fractions (fine coal separation). This Centrifugal Jig also combines.centrifugal acceleration (35 G) with jigging. The Altair Jig consists of a rotating bowl, which is placed inside a static casing having separate launders for collecting the concentrate and tailing samples. The rotating bowl contains a cylindrical screen with a lip, whose height can be adjusted to vary the natural depth of the ragging bed. The ragging material remains in a vertical position on the screen due to the rotation of the bowl.

In The Altair Jig, feed (slurry) which is introduced from the top at the centre of the rotating bowl is distributed into the ragging bed on the screen by the diffuser plate placed under the feed inlet. Pressurized water is injected under the bed periodically through the four pulse blocks to cause alternating dilation and contraction of the ragging and feed bed. This action, coupled with the high centrifugal force generated from the rotation of the bowl, provides the jigging mechanisms and kinetics required to achieve gravity-based separation for fine and ultrafine particles.

In case of fine coal separation, the tailings, which contain particles having a relatively high density, settle through the ragging bed and screen into the hutch and reports to the tailings launder through the discharge ports. On the other hand, the clean coal particles, which have a relatively low density do not have sufficient retention time to settle through the screen, and, thus, report to the concentrate launder. Based on fine coal separation testwork (Pittsburg No. 8 seam coal sample), the identified key operating 12

parameters were the drum speed, pulse water pressure, ragging volume and volumetric feed flow rate (Mohanty et al, 1999).

2.4.1.4. Mozley Multi Gravity Separator (MGS)

The Mozley Multi Gravity Separator (MGS) was developed by Richard Mozley Ltd. to improve the economic recovery of very fine heavy minerals and utilizes centrifugal action to increase the gravitational force acting on mineral particles. The MGS combines centrifugal acceleration (8 to 22 G) with the motion of a shaking table to achieve separation. The MGS is effective at recovering particle sizes ranging from 300 Lim to 10 |im and can produce high enrichment ratios even at fine particle sizes.

The principle of the MGS may be visualized as rolling the horizontal surface of a conventional shaking table into a drum and rotating it. The unit consists of two identical drums rotating in opposite directions to maintain mechanical equilibrium. Feed to both drums is gravity fed from a splitter box arrangement. Both the wash water and feed are directed onto separate accelerator rings located within the drum which reduces the velocity and prevents disturbance of the particle bed. Feed enters the drum about half way and wash water is supplied by a perforated ring near the open end of the drum. The drum motion causes segregation of high density particles through to the inner wall of the drum. The shaking motion disturbs the bed to allow high density particles to migrate to the wall. Scrapers moving slightly faster than the drum transport the concentrate counter to the flow toward the smaller end of the drum to a launder. The low density particles flow with water towards the larger end of the drum to a tailing launder.

Parameters affecting MGS performance are feed pulp density, drum speed, throughput, wash water flow rate, angle of tilt and shake intensity. The main short term control variable is the wash water flow rate (Turner et al., 1993). 13

2.4.1.5. Knelson CVD

The Knelson CVD is a continuous variable discharge concentrator for high mass yield applications. The machine operates on the same principle as the Knelson batch concentrators but draws the concentrate through pneumatically actuated pinch valves located radially in the rings.

In recent gravity separation developments, more difficult sulphide associated gold ores can be processed using the Knelson CVD. The CVD can capture both the fine free particles of gold and the gold associated sulphide particles. The CVD can be placed in various locations of the gold milling circuit. Specific sites may warrant a different placement as determined by mineralogy and site layout (Knelson, 2003).

In the CVD, feed is introduced to the top of the machine through a feed tube into the centre of the bowl section. The feed hits a plate at the bottom of the bowl section and is dispersed radially to the bowl wall. The particles are accelerated to a g-forced defined by the bowl speed and travel up the wall towards the ring. The upgraded slurry enters the separation ring where fluidization water, supplied through holes in the ring wall, is added to fluidize the bed of packed particles. Concentrate is extracted through pinch valves at the back of the ring. The pinch valve open and closed time can be adjusted. The light particles overflow the bowl into a tailings launder (Knelson, 2001).

The CVD has four main operating variables: bowl speed (g-force), fluidization water flow rate, pinch valve open time and pinch valve closed time. These variables all interact, meaning, for example that the best bowl speed at one set of pinch valve open/close times may not be the best at another set of times. Therefore, the appropriate selection of variable levels is a complicated problem that requires a systematic approach (Mcleavy et al, 2001). 14

2.4.2. Applications

There have been significant innovative advances for all types of gravity concentration equipment. The best application for these separators needs more study and is affected by unit capacity, inability to handle coarse particles, water requirements and metallurgical performance. Despite the present limitations, further development of continuous centrifugal concentrators has the potential to significantly change mineral processing plant flowsheet design. Applications and operating parameters of the several continuous centrifugal concentrators are listed in Table 2.1.

Table 2.1. Continuous Centrifugal Gravity Concentrators Technologies and Applications

Max. Mass G- Solid Particle Size Technology Capacity Yield Force Application (%) (mm) (t/h) (%) Range Primary recovery", Knelson CVD 100 25-55 0.1-40 -1 +0.025 60-120 Pre-concentration*, Scavenging" Primary recovery, Falcon Model C 100 25-45 0.1-50 -1 +0.025 50-300 Pre-concentration*, Scavenging* — 1 Primary recovery , Altair Centrifugal 18 25-40 1.5-100 -0.3+0.025 Secondary treatment, Jig 30-160 Unit 35 Scavenging", Pre- concentration* Primary recovery*, Secondary treatment, Kelsey Jig 65 25-65 10 -0.3+0.010 50 Scavenging, Pre- concentration*, Mozley C902 4 Primary recovery, Multi- MeGaSep 30 1-10 -0.3+0.010 8-22 Pre-concentration, 60 Gravity (Coal) Secondary treatment,

* Tested Application ** Potential Application 15

2.4.2.1. Falcon Model C

The Falcon Model C, with a capacity of up to 100 tph has potential applications for pre- concentration of heavy minerals and scavenging plant tailings. This machine is suitable for a wide range of minerals and fine feed applications requiring low to high weight recoveries (mass yield up to 50%) such as primary recovery of fine gold, fine coal cleaning, and fine iron ore and tin recovery. Applications of the several Falcon concentrators are listed in Table 2.2.

Table 2.2. Falcon Model C Operating Applications

Customer Country Model Units Application EQUIPROM Iron fines Brazil C-400 1 concentration plant Gold pre-concentration Newcrest Mining Australia C-4000 1 trial Tantalum recovery- Sons of Gwalia, C-4000 1 rougher Wodgina Tantalum Australia Tantalum recovery- Mine C-1000 1 cleaner Enviro Gold Canada C-1000 1 Gold recovery Evis House Inc., Canada C-400 Mill clean up Gold 1 Metals Research Canada C-1000 Primary gold recovery Corp. 1 Tantalum recovery- C-1000 1 rougher Tantalum recovery- Tantalum Mining C-400 1 cleaner Corp., TANCO Canada Spodumene recovery- Mine C-1000 1 rougher Spodumene recovery- C-400 I cleaner Silver recovery- C-1000 Silver Eagle rougher Resources, Silver Chile tailings treatment Silver recovery-cleaner C-400 • EMINZA Group Ecuador C-400 Gold Telluride of Companies 1 International Ghana C-400 Gold fines - Pilot plant Tournigan Corp. 1 DZAMGYR Kyrgyzstan C-400 1 Gold recovery Hematite recovery Mikhailovsky Gok Russia C-1000 from tailings ' (Scavenger) 16

Customer Country Model Units Application C-400 1 Metso Minerals Russia Gold recovery C-1000 1 Multotec Process South Africa C-1000 1 Pilot plant Flotation cone, Gulf International Tajikistan C-1000 1 cleaning Sulphide and electrum Echo Bay Minerals USA C-4000 1 recovery

A Model C-4000 was used to concentrate gold-sulphides at Echo Bays's Kettle

River Mine, Washington. The feed was overflow from the hydrocyclone (90 tph). The

Falcon concentrate is subjected to high intensity cyanidation and the tailings are leached in the pre-existing Carbon In Pulp (CIP) circuit. The operators reported an increase in production of approximately 3,000 ounces per year (increased overall gold recovery of

2.7%) while maintaining cyanide consumption at previous concentration level. Daily throughput was approximately 2,000 tons (Sprake et al, 2003).

Falcon installed two C-1000 and two C-400 units in the TANCO tantalite mine in

Manitoba, Canada in 1998. The tantalum ore is ground to -2 mm and concentrated using a combination of gravity separation methods (spirals, Falcon C, Mozley multi-gravity separators, cross-belt separator and shaking tables). Different circuits recover different size fractions of the tantalum, for an overall recovery in the 70% range. The tantalum concentrate typically grades between 35% and 38% Ta2C»5, with smaller percentage of tin, niobium, and titanium (Hilliard, 2003). A typical installation of two Falcon C continuous concentrators in series treating up to 100 ton per hour of -1mm material

(tailings from the spirals and tables in TANCO Mine) and the second unit as the primary concentrate cleaner. The effective use of this circuit yields high recoveries of the target mineral/metal in a fraction of the feed mass, producing a low tonnage stream for the final table cleaning unit (Falcon website, 2004). 17

2.4.2.2. Kelsey Jig

The high efficiency of separation, as well as the introduction of the high-capacity of the

Kelsey Jig, makes it an excellent option for processing existing tailing deposits and scavenging current plant tailing streams to recover fine valuable minerals, as well as for removing environmentally unacceptable minerals from plant tailings (Richards et al,

2004).

The Kelsey Jig is used to recover cassiterite, tantalum and chromite. There are relatively few gold installations, but it:

Has demonstrated the ability to recover fine flaky placer gold;

Is used to scavenge fine gold from shaking table tailings (secondary treatment);

Has been tested for the recovery of gold from the cyclone overflow, where it

recovered gold as fine as 5 um.

Applications of the several Kelsey Jig concentrators are listed in Table 2.3.

Table 2.3. The Kelsey Jig Operating Applications

Customer Country Model Units Application Bakrychik Gold Kazakhstan J650 1 Gold/sulfides (Scavenger)

Gwalia Consolidated Australia J1300, J650 2 Tantalum recovery-rougher

Lonrho Platinum South Africa Chromite Beijing Mining Inst. China Inbras Equpamentos Paranapanema Brazil Tin recovery J1300MKII 1 Mineracao Renison Tin Ltd. Australia J1300, J650 7 BHP Minerals Cable Sands Australia J1300 1 CRA Ltd. \ Namakwa South Africa Mineral sands Richards Bay Minerals RGC Mineral Sands Australia/USA TiWestJV Australia J1300 2 Comalco Research Australia Bauxite Du Pont Corp USA J1800 2 Titanium recovery 18

Customer Country Model Units Application Magma Minerals USA PGM's Regional Research Lab India Iron ore SINTEF Norway Iron ore/titanium Comsur Bolivia J1800 1 Placer Dome Asia Australia J1800 3 WMC Australia J1800 1 Minsur Peru J1300 MKII 1

A Model J650 is in production at the Bakrychik gold mine in Kazakhstan. The

unit improves the overall gold recovery by scavenging fine gold-sulfide particles from the

table circuit tails. The concentrate is reground and returned to the head of the circuit. A

major portion of gold recovery (56.7%) is in the fine fraction demonstrating the ability of

the Kelsey Jig to recover finer particles than conventional concentrators such as shaking

tables (Jones, 1999).

At the Gwalia Consolidated tantalum mine, Australia, the Kelsey Jig achieved an

upgrade ratio of 80 times at a 68% recovery for ore ground to 50% -75 um (Clifford,

1999). However; testing to separate ultrafine (45% -12 pm) tantalum minerals (S.G. 6.5-

7.2) from slimes yielded poor results (Burt et al, 1995).

2.4.2.3. Altair Centrifugal Jig

There are no references for commercial operations using the Altair Jig. A 30-16 (screen

diameter - screen height in inches) unit was installed for testing at a plant to separate

zircon from alumina-silicate. Limited testing has been conducted aimed at recovering

gold; however it has demonstrated the ability to recover fine flaky gold and fine coal.

The capacity of the 30-16 unit is 18 tph and for 83-16 unit is 38 tph.

A pilot scale study conducted on -600 + 44 pm coal indicated that the centrifugal jig is capable of separating materials at low densities and with small density differentials

over a range of fine particle sizes (Mohanty et al, 1999). The probable error (Ep) ranged

from 0.08 at a separation density of 1.54 to 0.17 at a separation density of 1.43. For

particles larger than 300 pm, selectivity was found to decrease, which was explained by

the restricted flow of the particles through the screen. Good separation efficiency was 19 demonstrated for particles as fine as 44 um, although the maximum separation efficiency was achieved for the -150 + 75 urn size range.

The Altair Jig has the potential application to separate efficiently at fine particle sizes (300 um to 25 um), at low density differences with product yield from 0 to 100% as required. Limitations are low capacity and high water requirement. Table 2.4 presents the tested applications for the Altair Centrifugal Jig.

Table 2.4. Tested applications for the Altair Centrifugal Jig

Heavy/SG Zircon/4.68 Hematite/5.26 Rutile/4.2 Light/SG Kyanite/3.5 Quartz/2.65 Quartz/2.65 Concentration Criterion 1.41 2.56 1.94 %-200 mesh Not specified 29.2 78.8 Weight % 14.6 10.8 28.8 Grade % 37.4 63.6 37.3 Recovery % 80.3 59.4 85.8

For gold processing, tests have been conducted to scavenge gold from secondary treatment tailings (shaking table and Knelson tails).

2.4.2.4. Mozley Multi Gravity Separator (MGS)

The MGS is effective for separation of fine and ultrafine particles. Laplante et al (1998) conducted tests on samples containing gold from a flash flotation concentrate diluted with quartz. The MGS was found to be very effective at recovering a full range of gold particle sizes from 300 um to 10 um and out performed a Knelson at the fine sizes.

Laplante stated that the MGS can not compete with the Knelson for the recovery of gold from grinding circuits due to its low capacity and high water usage. However, he suggested that its ability to recover very fine gold effectively may be applied to recovering gold from base metal concentrates with low gold smelter returns such as zinc concentrates. The high enrichment ratio achieved at fine particle sizes indicates the potential application for scavenging fine gold from shaking table (secondary treatment) tailings. 20

The MGS has been applied at the flotation circuit of the Wheal Jane mill,

Cornwall, UK in 1989. The goal was to improve the tin concentrate grade and plant metallurgical performance by replacing column flotation with a MGS, as the final stage of cleaning (Turner et al., 1993). Tests with tantalum slimes (45% -12 pm) indicated far superior enrichment compared to the Falcon, Knelson and the Kelsey Jig. Open circuit treatment of chromite slimes from an Albanian deposit grading 23% Cr203 produced a concentrate grading 50% Cr203 with a recovery of 72%. It has also demonstrated good separation of fine pyrite from coal, fine tungsten and base metal flotation concentrates.

The MGS has also been applied to wolframite recovery from plant tailings (as a scavenger) at mineral Regina, Peru and upgrading graphite flotation concentrates in

South Australia. The low capacity of the MGS limits its potential application (Model

C902 - 4 tph), although a higher capacity unit has been built for coal preparation

(MeGaSep - 60 tph).

In TANCO Mine the tantalum ore is ground to -2 mm and concentrated using a combination of gravity separation methods such as spirals, Falcon C, MGS, cross-belt separator and shaking tables (Hilliard, 2003). 21

2.4.2.5. Knelson CVD

Preliminary tests were conducted on synthetic mixtures of magnetite (density 5 similar to

pyrite) and quartz ground to approximately 80% passing 100 mesh at the University of

British Columbia. A single stage of processing increased the magnetite grade from 4% to

18.1% recovering 82.5% of the magnetite in 19.3% of the feed weight (reject 80.7%).

Rougher-scavenger configurations are being evaluated to improve recoveries further

(Lambert et al, 1999).

Most of the CVD applications involved pilot scale testing in the laboratory or on

mine site, using a 2 tph CVD6. There are CVD 20 and CVD32 units that were installed

in Russia and Africa, but results are scant and optimization is as yet incomplete

(Simpson, 2003).

Table 2.5. The Knelson CVD Tested and Operating Applications

Country Model Units Application Mali CVD6 1 Pilot test, gold-sulfides

Ghana CVD6 1 Pilot test, gold-sulfides

Papua New Guinea CVD6 1 Pilot test, gold-sulfides Ghana (Bogoso) CVD32 1 Gold-sulfides Russia (Turbokon) CVD32 1 Tin recovery Russia (Zapadnaya) CVD20 2 Gold-sulfides Zimbabwe CVD20 1 Alluvial chromite Canada (Luzenac) CVD32 1 Talc cleaning-reverse separation

There is a CVD32 that has been installed in Luzenac talc mine in Ontario,

Canada. Iron compounds such as magnetite, goethite and hematite are commonly found

as minor contaminants in Luzenac talc mine. Iron compounds are often entrained in froth

flotation concentrate and require removal prior to shipment.

In Luzenac beneficiation plant CVD is typically incorporated into the flotation

circuit treating flotation concentrates. The first cleaner concentrate is fed to the CVD, producing light and heavy product streams. The light stream of the CVD has most of the 22 iron compounds removed, and reports to the second cleaning steps. The rejects stream from the CVD containing both fine liberated iron particles and coarse talc particles with locked iron are reground and sent back to primary flotation (Byron et al., 2004).

2.5. Conclusion

Metal values in coarse particle fractions represent a considerable amount of the total metal loss by most base metal flotation operations. The potential to improve this recovery by even a small amount will have enormous benefits to the mining industry.

The continuous centrifugal gravity separators represent a relatively new technology that were developed and manufactured by above mentioned companies.

There is a top limit for particle size floatability which varies for different ores.

The coarser fractions in flotation feeds contain heavy and valuable liberated and non- liberated minerals (middling) that cannot be floated efficiently. The specific gravity of these particles plays an important role which causes decreasing flotation performance.

Due to their relative density differences, middling particles can be separated from gangue using gravity-based processes. For treating fine particles, the ability of gravity separators has been limited due to lack of particle inertia relative to the surface drag forces.

However, particle inertia can be enhanced by the application of a centrifugal field.

The size enhanced density separator used in this investigation was a Knelson

CVD having a 6 inch diameter bowl. Application of this new technology in scavenging tailings has the potential to improve metal recovery and reduce grinding power requirements and associated costs. This research is aimed at assessing the performance of the CVD for this application. 23

CHAPTER 3 Experimental Program

3.1. Analyzed Samples

3.1.1. Eskay Creek Samples

Two samples (600 kg solid of each) referred to as Low Grade Tailings (LGT) and High

Grade Tailings (HGT) were obtained from the beneficiation plant of the Eskay Creek

Mine, Barrick Gold Corp. Figure 3.1 shows the sample point of the beneficiation plant circuit.

Tailings Feed Bail Mill Hydrocyclone Flotation 1 Cyclone w O/F Cone.

Batch Knelson

Cone. r Gemini Table

Cone. r Smelter

Figure 3.1. Eskay Creek Flotation Circuit Sample Point 24

3.1.2. INCO Thompson Samples

Two samples (300 kg solid of each) referred to as Scavenger Tailings (AS-4) and

Scavenger Cleaner Tailings (SC-4) were obtained from the beneficiation plant of the

Thompson Mine, INCO Manitoba Division. Figure 3.2 shows the sample points of the

beneficiation plant circuit.

AS-4 Jl££|U Roughers Regrind Scavengers Rougher-Scavenger Tails Rougher Con

Rougher-Cleaner Tails SC-4 U Rghr-Clnrs OH Scav-Clnrs Scavenger-Cleaner Tails Xu/Ni Con Cu/Ni Tails Smelter Ni Con Cu/Ni Sep.

Cu-Cleaners Tails J>| Cu-Clnrs

Smelter Cu Con

Figure 3.2. Birchtree Flotation Circuit Sample Points 25

3.2. Sample Characterization

3.2.1. Sample Preparation

Representative 40 kg sub-samples were obtained from each wet tailing sample. The

sampling facility consisted of an agitated tank and slurry feed pump. Each sample was

fed to the tank and agitated for about two hours. Then the circulation process was begun

in a closed circuit for about six minutes to ensure steady-state conditions prior to

sampling. To obtain the required sub-samples, the slurry was sampled every two minutes

for ten seconds from the agitated tank discharging pipe. Figure 3.3 shows the sampling

procedure block diagram.

The collected samples were decanted and dried. The dried samples were riffled to

obtain representative portions for:

Size-assay analysis Density fractionation of sized fractions for elemental characterization

Scoping gravity recovery test using a batch 3 inch Knelson.

Tailing Sample

Pump Box Split Sample Feed Tank

Decantation and Drying

Riffling Assay

Screen analysis

LKC T Magstream IE Assay Concentrate Assay Tailings Assay \4

Figure 3.3. Sampling procedure of wet tailing samples 26

3.2.2. Size/Assay Analysis

The portions were screened using US Sieve Numbers 50, 70, 100, 140, 200, 270 and 400

and the size fractions were weighed. Each fraction was subjected to multi-element ICP

analysis to obtain grades of Ni, Cu, Fe, Mg, Pb, As, S, Ag and Au (All assays were

conducted by ACME Analytical Laboratories, Vancouver).

3.2.3. Mineralogical Characterization

The representative portion of the Eskay Creek tailing sample (LGT) was screened using

US Sieve Numbers 100, 200 and 400 and the size fractions were weighed. Polished

sections of the each size fraction were examined using optical microscopy. The

mineralogical work was focused on identifying the occurrence, size, association and

degree of liberation of the gold-sulphide grains. This part of work has been performed by

undergraduate students during their process mineralogy course (Mine 338) under

supervision of Dr. Klein.

3.2.4. Density Fractionation for INCO Samples

Prior to density fractionation, a hand magnet was used to separate the ferromagnetic

particles. A Magstream was then used for separation at densities of 2.7, 3.1, 3.5, 3.9 and

4.3 g cm"3. The removed magnetic fractions of each size fraction were then added to the

density fraction with density grater than 4.3 to determine mass and metal distribution as

they are mostly pyrrhotite. The density fractions were weighed and assayed. The data

were then used for calculating the partition coefficient numbers.

The Magstream process in general can be described as a combination between magnetic and gravimetric separation. It is used for non-magnetic or weakly magnetic particulate materials. Operation is similar to that of a heavy media centrifuge, except that that the effective density of the liquid is magnetically derived, allowing it to be adjusted 27 without practical limit by changing the speed of rotation or concentration (magnetic strenghth) of the process fluid (IGC, 1991). The sample is in slurry with the process ferrofluid (magnetite in water) and then it enters through a long rotating shaft. The material that passes through gets separated into a low density and high density stream.

There are several advantages of this process in comparison to conventional heavy-media sink-float processes. The cost and the toxicity of heavy liquids make the Magstream more appealing to use, also it is much safer for the operator because no fumes are being produced. The Magstream process can be used to separate material at much higher densities to over 8, while most heavy liquids do not go to over density of 5.1 g cm"3. 28

3.3. Knelson Batch Tests

For the gravity concentration scoping test, a 2kg portion of each sample was fed to a 3 inch laboratory Knelson Concentrator (LKC). The products were weighed and assayed for metallurgical balances. Figure 3.4 shows the block diagram of the batch tests.

Head Sample

Pan Con. Pan I 1 LKC 1 Con.

Pan Tail. 5 Pan Con. 2 LKC 2 Pan 2

Concentrate

Pan Con. 3 LKC 3 Pan 3

Pan Con. 4 LKC 4 Pan 4

Jt±=z Tailings

Figure 3.4. Batch Knelson test procedure 29

3.4. Knelson CVD Pilot Scale Tests

Based on the sample characterization and batch scale test results, pilot scale tests (seven tests for SC-4 and six for LGT) were conducted. The pilot scale testing facilities consist of an agitated feed tank, a CVD6 concentrator and a Sala pump that pumps the combined concentrate and tailing products to a storage tank. The glass beads (total weight: 309 gram, bead diameter: 6 mm and density: 2.7 g cm'3) were used for ragging. The test operating conditions are summarized in Tables 3.1 and 3.2.

To prepare feed for pilot testing, the feed tank agitator was turned on for about two hours. The pulp density was adjusted by adding water to 25% solid. During the first

6 minutes of the test operation no sampling was performed to ensure steady-state operating condition. For the first 6 tests of SC-4 and the first five tests of LGT the concentrates were sampled for 6 minutes (from time six to twelve minutes), and the tailing product was sampled at 7, 9 and 11 minutes of operation for 10 seconds each. The sampling times and number of cuts were recorded to calculate the mass yields. After drying and weighing, the products were riffled to obtain sub-samples for size/assay analysis. The left over material was returned to the feed tank. For the last pilot test (test

7 for SC-4 and test 6 for LGT), the concentrate sampling was started after 6 minutes and continued until feed tank was empty. For this test the tailing product was sampled at 7,

11, 15, 19, 23 and 27 minutes of operation for 10 seconds of each sampling time.

Table 3.1. Test Operating Conditions, LGT

Flow Rate Fluidization Pinch open Pinch closed Bowl Speed Test No. (lit/min) Water (gpm) Time (s) Time (s) (rpm)

I 43 8 0.03 0.8 830 2 43 8 0.03 0.5 830 3 43 8 0.03 0.2 830 4 43 8 0.03 0.2 650 5 44 8 0.03 0.2 1050 6 42 8 0.03 0.2 830 30

Table 3.2. Test Operating Conditions, SC-4

Flow Rate Fluidization Pinch Open Pinch Closed Bowl Speed Test No. (lit/min) Water (gpm) Time (s) Time (s) (rpm)

1 42 8 0.03 0.8 .830 2 50 8 0.03 0.8 830 3 42 8 0.03 0.5 830 4 43 8 0.03 1.1 830 5a 42 8 0.03 0.8 500 5b 42 8 0.03 0.8 500 6 42 8 0.03 0.5 500 7 35 8 0.03 0.5 830

3.4.1. Products Characterization

The representative portion of all concentrate and tailing products were screened using US

Sieve Numbers 50, 70, 100, 140, 200, 270 and 400 and the size fractions were weighed.

Each fraction was subjected to multi-element ICP analysis to obtain grades of Ni, Cu, Fe,

Mg, Pb, As, S, Ag and Au (All assays were conducted by ACME Analytical

Laboratories, Vancouver). The high Ni results were re-assayed using an Aqua Regia

Digestion method followed by ICP. The Au and Ag results were confirmed using fire assay.

Metallurgical balance was prepared based on measured mass flow rates and grades of concentrates and tails. These calculated mass yields were reasonable with respect to expected changes resulting from changes in operating variables. Average calculated feed grades were comparable to measured feed grades. Mass yields based on grades of feed, concentrates and tails were variable due to low mass yield in concentrate and finer tail grades were similar to feed grades (i.e. mass yield calculations were very sensitive to small variation in tails grades).

Polished sections of the each size fractions of the obtained products of the pilot testing (Test 6, Eskay Creek LGT Sample) were used to examine the occurrence, size, association and degree of liberation of the gold-sulphide grains using optical microscopy. 31

3.4.2. Concentrate Density Fractionation Tests

Prior to density fractionation test, a hand magnet was used to separate the ferromagnetic particles from the SC-4 concentrate of Test 7. This product was screened using the US

Sieve Numbers: 50, 100, 140, 200 and the size fractions were used for density fractionation at densities of 2.7, 3.1, 3.5, 3.9 and 4.3 g cm"3 using a Magstream. The removed magnetic fractions of each size fraction were added to the density fraction with density grater than 4.3 to determine mass and metal distribution as they are mostly pyrrhotite. The density fractions were weighed and assayed in order to find the partition coefficient numbers and cut point.

3.4.3. Flotation Test and Proposed Flowsheet

Flotation tests were conducted on the CVD concentrates of both LGT and SC-4 samples.

The tests were performed based on the existing plant conditions of grinding and flotation circuits. The test conditions and metallurgical balances are shown in Appendix E. 32

CHAPTER 4 Results and Discussions

4.1. Eskay Creek Samples

4.1.1. Introduction

The Eskay Creek Mine comprises of different ore zones. Table 4.1 shows the typical ore grades for two main zones.

Table 4.1. Ore Zones Assays (1998)

Ore Zone Au (g/t) Ag (g/t) Cu% Pb% Zn% 109 27.4 76.5 0.02 1.03 1.73 NEX 101.0 2026.0 0.9 3.82 6.02

Based on processing plant results from 1998, the cyclone overflow with a P80 of

74 reports to the flotation feed conditioner. Depending on the ore type, the total gold recovery by gravity and flotation ranged from 94 to 97%. The GRG (Gravity

Recoverable Gold) depends on the ore type and ranges from 30 to 45%. The flotation silver recovery ranged from 90 to 95%. For base metals such as copper, lead and zinc the recoveries are also ore type dependant and all are expected to range from 85 to 95%

(Anand et al, 1998).

The present practice involves mixing of different ore types. The plant was sampled in May, 2003 and showed that the gold recovery has decreased to 92.9%. These results are based on four different types of ore (NEX, 21B, 109 and 21C) that were fed to the processing plant. The specification of beneficiation plant feed and two received tailing samples are shown in Table 4.2. 33

Table 4.2. Ore Types Percentages (Plant Feed) and Tailing Sample grades (Data from

Eskay Creek Plant, 2003)

Ore Types % (plant Feed) Au Ag Tailing Grind Size NEX 21 B 109 21 C Total Samples ppm ppm % -75 pm 36 38 19 7 100 70.3 LGT 1.9 48.3 49 24 12 15 100 72.6 HGT 3.2 39.0

Tables A. 1.3 and A. 1.4 (Appendix A) present the size/assay results and also the mill performance for both received samples.

4.1.2. Results and Discussions

4.1.2.1. Sample Characterization

4.1.2.1.1. Size/Assay Analysis

The size/assay analysis of Eskay Creek tailings revealed that the Au grade for LGT and

HGT are 1.1 and 1.9 ppm, respectively. The Pso for LGT sample is 130 pm and for HGT is 100 pm. The +100 mesh size fraction contains about 9% of mass of LGT and about

5% of HGT. The gold distributions in this size fraction are 8.9% and 3.3%, respectively.

The respective sulphur distributions are 7.0% and 3.4%.

Figures 4.1 and 4.2 show the gold and sulphur grade distributions in each size fraction of both samples. Based on these data, the following results can be obtained:

1. The gold and sulphur grade distributions shows that gold and sulphur

losses occur at fine (-37 pm) and coarse (+210 pm) fractions.

2. Grinding coarser would reduce losses in the fine fraction.

The objective of the CVD testing is to recover free gold and gold-sulphides from the coarsest fraction. 34

2.5

• LGT H HGT 2.0

E Q. 1.5 Q. - e> • •o % 2 (3 1.0 •A a • <

"la 0.5 il.

I • 0.0 -37 -53437 -74+53 -105+74 -149+105 -210+149 -297+210 +297

Size Class (um)

Figure 4.1. Au Grade Distributions, LGT and HGT

2.0

• LGT D HGT

1.5

0> | 1.0 o

0.5

0.0 -37 -53+37 -74+53 -105+74 -149+105-210+149-297+210 +297

Size Class (um)

Figure 4.2. S Grade Distributions, LGT and HGT 35

It is believed that the CVD will recover particles based primarily on density but also based on size. Mass yields, metal recoveries and grade targets are arbitrarily set as the mass yield in the +149 pm fraction, Au recovery of the +105 pm fraction was taken as the target and grades were calculated from this fraction. Table 4.3 summarizes the results.

Table 4.3. Samples Characterization Results, Target Mass Yields and Metal Recoveries

+149 pm Fraction +105 pm Fraction Sample Pso (pm) Weight % Au Distribution% S Distribution% LGT 130 9.3 27.9 20.6 HGT 100 5.3 10.6 10.7

For the LGT sample, the target mass yield is 9.3%, Au recovery is 27.9% and grade is 3.4 ppm. For the HGT sample, a mass yield of 5.3%, Au recovery of 10.6% and

Au grade of 3.8 ppm are targeted. These numbers indicate that both samples have potential for middling recovery. In both cases, the gold distribution is greater than the mass distribution indicating the potential to upgrade the gold bearing minerals.

The results of both the sample characterization and size/assay analysis are shown in Appendix A.

4.1.2.1.2. Mineralogical Characterization

Optical microscopy revealed that in all size fractions above 37 (im, sulphide minerals including pyrite and arsenopyrite, are primarily associated with gangue minerals

(silicates) as fine inclusions. The degree of liberation for gold bearing sulphide grains of each size fraction of the sample were examined applying the point counting method. The grains were counted and split into five sections. The degree of liberation is calculated based on the percentage of the free particles divided by the sum of the free and locked particles. Table 4.4 shows the calculation results for LGT sample. The results indicate that there are no liberated sulphide grains coarser than 150 |im. 36

Table 4.4. Degree of Liberation Results, LGT

Size Fraction (|im) Degree of Liberation % +149 0 -149 + 74 50 -74 + 37 52 -37 82

The mineralogical analysis results of each size fraction of the sample are shown in

Appendix D.

4.1.2.2. Batch Tests Results

The results of the batch Knelson tests are summarized in Table 4.5, which shows the

weight %, Au grade and Au distribution for the concentrates.

Table 4.5. Batch Test Results

Cone. Weight Au Grade Sample Au Distribution % Recovery % (ppm) LGT 5.0 5.1 20.0 HGT 4.0 3.9 8.8

The metallurgical balances are presented in Appendix B. For the HGT, results are

similar to those predicted from the characterization results. For the LGT sample, while

the weight and Au distributions are lower than the target value, the Au grade is higher.

4.1.2.3. Pilot Scale Tests Results

Based on the characterization and batch Knelson tests, there appeared to be a greater

opportunity to recover a recyclable product from the LGT sample than the HGT sample.

It was therefore selected for pilot scale tests using the CVD6 separator. These tests were

conducted with different operating variables including pinch valve closed time and bowl

speed. These plant testing produced a range of results. The pinch open time was maintained at 0.03 seconds and fluidization water flow rate maintained at 8 gallon per 37 minute for all tests. As shown in Table 4.6, the best result was obtained from Test 3 based on pinch valve closed time of 0.2 seconds and bowl speed of 830 rpm (60G).

The pilot scale test results were used to evaluate the possibility of improving gold recovery from the LGT sample of the Eskay Creek beneficiation plant. These results were also used to evaluate the Knelson CVD concentrator as a scavenger for coarse middling particles from flotation tailing and to identify the target mass yields and set points for the operating variables.

4.1.2.3.1. Products Characterization

Products from the pilot scale tests were subjected to size/assay and mineralogical analysis. The metallurgical balances for the all tests are summarized in Table 4.6.

Table 4.6. Pilot Scale Test Results, LGT

Cone. Grade Distribution % Test Mass No. Yield Au(ppm) S% Ag(ppm) As(ppm) Au S Ag As % 1 0.9 6.31 1.29 87.3 202.8 5.5 1.3 1.5 1.4 2 1.7 4.46 1.07 84.8 192.1 6.5 1.7 2.8 2.1 3 3.3 12.51 1.14 76.4 198.5 28.4 3.9 4.9 4.0 4 4.4 3.66 0.71 71.1 129.3 14.7 4.4 5.9 3.4 5 7.7 5.60 1.04 63.1 165.7 35.0 11.5 9.4 9.3 6 9.3 4.36 0.99 68.5 157.6 15.3 12.4 12.3 10.0

The results show that the gold distributions are significantly higher than sulphur and arsenic distributions, which indicates that free gold or gold associated with poorly floated minerals is being recovered. The best result was obtained from Test 3 in which

28% of the gold was recovered in a concentrate grading 12.5 ppm Au representing 3.3% of the feed mass. This represents a gold upgrade ratio of 8.7.

Figures 4.3 and 4.4 show the gold and sulphur grade distributions in each size fraction of concentrate and tailings of Test 3. Based on these data, the following results can be obtained: 38

1. The gold and sulphur grade in the coarse fraction of the concentrate

are higher than that for fine fractions

2. The gold grade in tailing is very low.

3. CVD has successfully upgraded the gold in the coarsest fraction of the

concentrate.

Grinding coarser would reduce losses in finer fractions and resulting in a greater portion of coarse middlings to report to the flotation tailings. These types of particles can be recovered using CVD and recycled to the grinding circuit. The overall consequence will improve overall gold recovery and reduce energy costs.

140 T ,

-37 -53+37 -74+53 -105+74 -149+105 -210+149-297+210 +297

Size Class (pm)

Figure 4.3. Au Grade versus Particle Size in Concentrate and Tailing of Test 3 39

I Concentrate. • Tailing

I ini -53+37 -74+53 -105+74 -149+105 -210+149 -297+210 +297 Size Class (um)

Figure 4.4. S Grade versus Particle Size in Concentrate and Tailing of Test 3

The test results show that the Au grade increased in the concentrate products significantly. This is a good indication that the centrifugal gravity concentrator can capture the middlings and the fine free particles of gold as well. The S grade increased in the finer fractions of the concentrate due to overgrinding soft sulphides. In case of Ag the results indicate no remarkable improvement. The table of metallurgical balances for all tests is shown in Appendix C.

Optical microscopy revealed that in all size fractions above 37 (Jm for CVD concentrate, sulphide minerals including pyrite and arsenopyrite, are primarily associated with gangue minerals (silicates) as fine inclusions. The mineralogical analysis results of each size fraction of the products are shown in Appendix D. 40

4.1.2.3.2. CVD Operating Conditions

Figure 4.5 shows the relationships between the mass yield, Au recovery and pinch valve closed time. The mass pull and the Au recovery decrease with increasing the pinch valve closed time. Figure 4.6 shows that the Au grade has a maximum value at the mass yield

3.5% and Au recovery increases with increasing the mass yield. The Au upgrade ratio has a maximum value at 3.5% concentrate mass yield (Figure 4.7).

Based on the metallurgical balance results of the conducted pilot scale tests, it is suggested to apply the operating variables as following: Pinch valve closed time of 0.2 seconds, pinch valve open time of 0.03 seconds, bowl speed of 830 rpm (60G) and water fluidization flow rate of 8 US gallon per minute (1.8 m3/h).

Figure 4.5. Mass Yield and Au Recovery versus Pinch Valve Closed Time Figure 4.7. Au Upgrade Ratio/Recovery versus Mass Yield 42

4.1.2.3.3. Size Partition

The size partition curves were obtained based on the results of all pilot scale tests. The feed mass flow rate of each test was calculated based on measured mass flow rate of concentrate and tailing products. These data and the size analysis results were then used for calculating the weight percentages and the mass flow rates for each size fraction (i) of concentrate, tailing and feed (Appendix C.3). The partition number (PN) for each size fraction was determined by using the following formula:

PN i = 100 * (concentrate mass flow rate j/feed mass flow rate;)

Figure 4.8 shows the size separation partition curves, which demonstrated that the

CVD operates as a size classifier as well as a density separator (size enhanced density separation). For Tests 3 and 6, the cut size was about 300 microns. The gold recovered was primarily in the coarse size fractions, which based on mineralogical analysis would likely contain middling particles (sulfides + silicates).

100 1000 Size (um)

Figure 4.8. Size Separation Partition Curves 43

Figure 4.9 presents the relationship between gold recovery and particle size. It shows that the gold recovery increases with increasing particle size. The separation efficiency performance is being affected by the feed characterization. The relationship between sulphur recovery and particle size is shown in Figure 4.10. The coarser the particle, the higher the sulphur recovery in concentrates. The increasing rate of sulphur recovery is lower than that for gold. The results of size/assay analysis for all CVD test products are presented in Appendix C.

100 1000 Size (Mm)

Figure 4.9. Au Recovery versus Size 44

80 I ; ! i 70 i l I i -A- - Test 3 ------Test 6 ' i 60 ! i I j 0^ 50

> 40 8$ rr 30 (0 20

10

0 10 100 1000 Size (Mm)

Figure 4.10. S Recovery versus Size

Figure 4.11 presents the relationship between gold recovery and sulphur recovery for CVD pilot tests (Tests 3 and 6). It shows that the gold recovery has a good correlation with sulphur recovery. Therefore, CVD could recover sulphides containing gold. The figures of all test results are presented in Appendix C.

S Recovery % Figure 4.11. Au Recovery versus S Recovery 45

4.1.2.4. Flotation Test Result and Proposed Flowsheet

The representative portion of CVD concentrate product (Pso = 180 pm) obtained from

Tests 6 of LGT sample was ground to 80% less than 75 pm (based on data from existing plant). The flotation test was conducted on the ground sample at natural pH and addition of PAX (100 g/t) and MIBC. The flotation test conditions and results are shown in

Appendix E. The aim of the test was to float the sulphides, which tends to maximize the recovery of the precious metals. The metallurgical balance is shown in Table 4.7.

Table 4.7. Flotation Test Results, LGT

Weight Grade Distribution % Material Au Ag Fe S % Au Ag Fe S (ppm) (ppm) % % Concentrate 33.4 9.26 141.0 3.85 2.3 81.0 72.9 37.6 74.3 Tailings 66.6 1.09 26.3 3.2 0.4 19.0 27.1 62.4 25.7 Feed (CVD Con) 100 3.82 64.6 3.42 1.03 100 100 100 100

The gold, silver and sulphur recoveries of the flotation test are high, which can have an impact on overall recovery of the beneficiation plant. Based on the existing plant results (93% Au recovery), CVD Test 6 (15.3% gold recovery from tailings) and followed by flotation (gold recovery of 81%), the overall gold recovery will be increased by about 1%.

It should be kept in mind that the samples used in this study are lower grade than the typical plant tailings. Therefore, it is expected overall improvements would be greater than 1%. Plant trials are recommended to assess this more accurately. The proposed flowsheet is shown in Figure 4.12. Tailings Feed Ball Mill Hydrocyclone Flotation Cyclone O/F

Cone..

Knelson Batch Knelson CVD

Cone. Final Tails r Gemini Table

Cone.

Smelter

Figure 4.12. Proposed Flowsheet, Eskay Creek 47

4.1.3. Conclusion and Recommendation

Two samples of flotation tailings referred to as High Grade Tails (HGT) and low Grade

Tails (LGT) from the Eskay Creek beneficiation plant were studied. The test program

was directed into three stages including sample characterization, batch gravity and pilot

scale CVD testing.

The data from sample characterization and batch tests on samples revealed that

both samples have potential for middling recovery. The Pso for LGT sample is 130 pm

and for HGT is 100 pm. A batch 3 inch Knelson was used to obtain a preliminary

assessment of gold recovery. For LGT sample the test revealed a concentrate grading 5.1

ppm with 5% weight and 20% gold recovery. For HGT sample, results are 3.9 ppm at

4% mass yield and 8.8% gold recovery.

Six pilot scale tests were performed on the LGT sample using the CVD6

separator. Products were subjected to size/assay analysis to assess separation. The

results showed the possibility of producing a concentrate with a low mass yield and high

gold upgrade ratio.

Pilot scale testing indicated that the CVD was effective at recovering gold from

flotation tailings. The results indicated that the gold recovered was primarily in the

coarse size fractions, which would likely contain middling particles (sulfide + silicates)

i.e. size enhanced density separation. The best result (Test3) was 3.3 % mass yield with

12.5 ppm Au grade and 28% Au recovery. For this test the Au upgrade ratio was 8.7.

Comparison of sulphur, silver and gold distribution indicates that free gold is

being recovered. According to the results attained, the CVD machine can upgrade the

flotation tailings of the Eskay Creek gold processing plant with a reasonable recovery. It

seems that the CVD recovers not only the coarser grains of gold bearing sulphides but

also captures the ultrafine free gold particles. The separation efficiency performance is

being affected by the feed characterization. 48

Size partition curves were obtained for all tests, which demonstrated that the CVD operates as a size classifier as well as a density separator (size enhanced density separation). For Tests 3 and 6, the cut size was about 300 microns.

Flotation test on ground CVD concentrate of LGT sample showed that, with recycling the CVD concentrate to the grinding and flotation circuit, the overall gold recovery will be increased by about 1%.

Modeling of grinding, flotation and CVD circuits to assess potential improvements in overall gold recovery and assess the potential for increasing grind size to reduce losses in ultrafine fraction, reduce energy requirements and increase throughput is recommended.

Based on the tests results and also the low mass stream of tailings at the Eskay

Creek beneficiation plant, onsite installation of a Knelson CVD6 is recommended to perform plant-trial test. 49

4.2. INCO Thompson Samples

4.2.1. Introduction

The Birchtree ore consists of pyrrhotite (both monoclinic which is ferromagnetic and hexagonal which is paramagnetic), pentlandite and minor chalcopyrite. This ore is closely associated with significant ultramafic rocks mostly serpentinites (magnesium

bearing materials) as well as talc. Typical ore composition and grades are shown in

Tables 4.8 and 4.9.

Table 4.8. Birchtree Ore Mineralogy (Wyshynski, et al, 1999)

Mineral Percent

Pentlandite, (NiFe)9S8 7.0

Chalcopyrite, CuFeS2 0.4

Pyrrhotite, Fe8S9 22.0 Graphite 0.5 Other 70.0

Table 4.9. Birchtree Ore, Concentrate and Tail Grades (Drapac, 2004)

Grade % Material Ni MgO Fe Ore 1.54 20.1 20.6 Final Ni Concentrate 10.8 6.7 36.5 Scavenger-Cleaner Tails (SC-4) 0.57 19.8 22.4 Rougher-Scavenger Tails (AS-4) 0.25 22.8 16.7

The nickel losses in Thompson flotation plant is about 20% of the total nickel of

the beneficiation plant feed. About 50% of this amount reports to the rougher-cleaner

tailing and 50% to the scavenger-cleaner tailing. The complexity of the ore mineralogy

and poor liberation may cause the low recovery. Talc plays an important role in froth

flotation process due to its natural hydrophobicity. Talc rejection from Birchtree ore may

improve the overall plant recovery. 50

4.2.2. Results and Discussions

4.2.2.1. Size/Assay Analysis

The size/assay analysis of Birchtree tailing samples revealed that the Ni grade for SC-4 and AS-4 are 0.5 and 0.2%, respectively. The Pgo for SC-4 sample is 94 urn and for AS-4 is 140 um. The +100 mesh size fraction contains about 10% of mass of SC-4 and about

16% of AS-4. The nickel distributions in this size fraction are 8.7% and 14%, respectively. The respective sulphur distributions are 9.7% and 7.4%. The Mg grade for

SC-4 and AS-4 are 10.4% and 11.8%, respectively.

Figures 4.13, 4.14 and 4.15 show the nickel, magnesium and sulphur distributions in each size fraction of both samples. Based on these data, the following results can be obtained:

1. The element distribution shows that metal losses at fine (-37 um) and

coarse (+149 um) fractions.

2. Grinding coarser would reduce losses in fine fraction.

The objective of the CVD testing is to recover nickel sulphides from the coarsest fraction. 7 -53+37 -74+53 -105+74 -149+105 -210+149 -297+210 +297

Size Class (um)

Figure 4.13. Ni Grade for each Size Fraction, SC-4 & AS-4 52

-37 -53+37 -74+53 -105+74 -149+105 -210+149 -297+210 +297

Size Class (um)

Figure 4.14. Mg Grade for each Size Fraction, SC-4 & AS-4

-37 -53+37 -74+53 -105+74 -149+105 -210+149 -297+210 +297

Size Class (Mm)

Figure 4.15. S Grade for each Size Fraction, SC-4 & AS-4

The CVD recovers particles based primarily on density but also based on size.

Mass yields, metal recoveries and grade targets are arbitrarily set as the mass yield in the 53

+149 um fraction, Ni recovery of the +105 um fraction and grades calculated from these numbers. Table 4.10 summarizes the results.

Table 4.10. Target Mass Yields and Metal Recoveries

+ 149um Fraction + 105 um Fraction Sample Pgo (nm) Weight % Ni Distribution% S Distribution% SC-4 94 10.2 14.0 16.6 AS-4 140 16.8 28.1 21.0

For the SC-4 sample, target mass yield is 10.2%, Ni recovery is 14.0% and grade is 0.68%. For the AS-4 sample, a mass yield of 16.8%, Ni recovery of 28.1% and Ni grade of 0.37% are targeted. These numbers indicate that both samples have potential for middling recovery using CVD. In both cases, the metal distribution is greater than the mass distribution indicating the potential to upgrade the metal bearing minerals.

Density fractionations for both samples were conducted using the Magstream. It was necessary to remove the magnetic particles from each size fraction as these particles will not separate effectively due to the magnetic field in the Magstream. Assays of the magnetic and non-magnetic components show that the Ni grade in the magnetic product is only slightly higher than in the non-magnetic product. Therefore, magnetic separation is likely not suitable for scavenging. The density fractionation data of SC-4 was used for plotting density partition curves in conjunction with the data of density fractionation tests on each size fraction of pilot scale test (Test 7) concentrate.

The results of the sample characterization, magnetic particle distribution, size/assay results of magnetic and non magnetic products and density fractionation for both SC-4 and AS-4 head samples are shown in Appendix A.

4.2.2.2. Batch Tests Results

The results of the batch Knelson test are summarized in Table 4.11, which shows the weight %, Ni/Mg grades and distributions in the concentrates. 54

Table 4.11. Batch Test Results, Concentrates

Cone. Weight Ni Grade Mg Grade Ni Distribution Mg Distribution Sample Recovery % . % % % % SC-4 9.5 0.70 3.4 14.0 3.3 AS-4 9.9 0.38 5.0 18.1 4.1

The tables including the detailed results and metallurgical balances are presented in Appendix B. For the SC-4 sample, the results are similar to the target results based on the size-assay analysis. For the AS-4 sample, while the Ni grade is similar to the target value, the weight and Ni distributions are lower. For SC-4, only 3.3% of Mg reported to the concentrate product, demonstrating the ability of the continuous concentrator to reject talc.

4.2.2.3. Pilot Scale Tests Results

Based on the characterization and batch Knelson test results and due to lower mass stream of the scavenger-cleaner tailing of the plant, SC-4 sample was selected for pilot scale tests using the CVD6 separator. These tests were conducted with different operating variables including pinch valve closed time and bowl speed. These pilot scale testing produced a range of results. The pinch open time was maintained at 0.03 seconds and fluidization water flow rate maintained at 8 gallon per minute for all tests. The best result was obtained from Test 3 based on pinch closed time of 0.5 seconds and bowl speed of 830 rpm (60G).

The pilot scale test results were used to evaluate the possibility of improving nickel and gold recoveries from flotation scavenger-cleaner tailings of the Birchtree ore processed at the Thompson beneficiation plant. These results were also used to investigate the possibility of decreasing the magnesium content from the product and finally to evaluate the CVD concentrator as a scavenger for coarse middling particles from flotation tailings. 55

4.2.2.3.1. Products Characterization

Products from the pilot scale tests were subjected to size/assay analysis. The metallurgical balances for the all tests are summarized in Table 4.12.

Table 4.12. Pilot Scale Test Results, SC-4

Cone. Grade Distribution % Test Mass Ni S Cu Mg Fe Au No. Yield Ni S Cu Mg Fe Au % % % % % (ppm) % 1 4.2 0.99 7.50 0.05 1.30 39.9 2.53 9.0 6.6 5.5 0.6 8.3 24.5 2 3.7 0.96 8.96 0.05 1.67 38.6 2.73 8.2 6.6 5.1 0.7 6.9 23.8 3 6.7 0.85 7.67 0.05 2.27 37.1 1.43 12.9 10.2 8.7 1.7 11.8 67.3 4 3.6 0.99 9.04 0.05 1.73 39.5 2.13 8.3 6.1 5.3 0.7 6.6 12.6 5b 0.3 1.43 6.19 0.07 1.42 41.5 3.95 1.1 0.4 0.6 0.0 0.7 18.2 6 0.5 1.13 7.47 0.06 1.11 45.6 9.99 1.5 0.8 1.4 0.1 1.2 52.4 7 9.4 0.72 5.80 0.04 2.86 33.0 2.34 16.1 13.5 11.8 3.2 16.1 60.1

The results show that the gold distributions are significantly higher than nickel and sulphur distributions, which indicates that free gold is being recovered. The best result was obtained from Test 3 in which 12.9% of the nickel was recovered in a concentrate grading 0.85% Ni representing 6.7% of the feed mass. It is worth noting that the Mg grade is decreased dramatically in the concentrate products. The Mg distribution in concentrate was just 1.7%. This is a good indication that the centrifugal concentrator can also remove the magnesium bearing minerals such as talc. In case of gold remarkable results were produced and the metallurgical balances showed gold recoveries ranging from 12.58 to 67.26% and gold grades from 1.4 to 10.0 ppm.

The pilot scale testing indicated that the CVD was effective at recovering gold and nickel bearing sulfides from the flotation tailings. The nickel and gold recovered were primarily in the coarse size fractions, which would likely contain middling particles

(sulfides + silicates). The results showed that CVD is able to reject the Mg bearing minerals as well. 56

Figures 4.16, 4.17, 4.18 and 4.19 show the nickel, magnesium, sulphur and gold grade distributions in each size fraction of concentrate and tailings of Test 3. Based on these data, the following results can be obtained: 1. CVD has successfully upgraded the nickel, gold and sulphur in the coarsest fraction of the concentrate. 2. CVD has reduced the magnesium in the concentrate. Grinding coarser would reduce losses in finer fractions and result in a greater portion of coarse middlings to report to the flotation tailings. These types of particles can be recovered using CVD and recycled to the grinding circuit. The overall consequence will improve overall nickel and gold recoveries and reduce energy costs.

2

-37 -53+37 -74+53 -105+74 -149+105 -210+149 -297+210 +297

Size Class (um)

Figure 4.16. Ni Grade for each Size Fraction of Concentrate & Tailing, Test 3 57

l Concentrate • Tailing

-37 -53+37 -74+53 -105+74 -149+105-210+149 -297+210 +297

Size Class (um)

Figure 4.17. Mg Grade for each Size Fraction of Concentrate & Tailing of Test 3

10

-37 -53+37 -74+53 -105+74 -149+105 -210+149 -297+210 +297

Size Class (um)

Figure 4.18. S Grade for each Size Fraction of Concentrate & Tailing, Test 3 58

8

7

E Q. 5 - a

a 3 3 2

1

0

Size Class (um)

Figure 4.19. Au Grade for each Size Fraction of Concentrate & Tailing, Test 3

The test results showed that the Ni, Au and S grades increased and the Mg grade decreased in the concentrate products. The Ni and S upgrade ratios for the coarser fractions (+149 um) are higher than that for fine fractions. In case of the Mg, the smaller the size fraction, the higher the Mg reduction. The table of metallurgical balances for all tests is shown in Appendix C.

4.1.2.3.2. CVD Operating Conditions

Figure 4.20 shows the relationships between the concentrate mass yield, Ni grade/recovery, Mg grade in concentrate and the pinch valve closed time. The mass yield and the Ni recovery decrease with increasing the pinch valve closed time. With increasing the pinch valve closed time, the Ni and Mg grades do not have a considerable change.

Based on the metallurgical balance results of the conducted pilot scale tests, it is suggested to apply the operating variables as following: Pinch valve closed time of 0.5 59 seconds, pinch valve open time of 0.03 seconds, bowl speed of 830 rpm (60G) and water fluidization flow rate of 8 gallon per minute.

0.4 0.5 0.6 0.7 0.8 0.9 1 1.1 Pinch Valve Closed Time (seconds)

Figure 4.20. Mass Yield, Ni Grade/Recovery versus Pinch Valve Closed Time

Figure 4.21 shows that the lower the concentrate mass yield, the higher the Ni upgrade ratio (Ni grade in concentrate/Ni grade in feed) and the higher the Mg downgrade ratio. 10 10

• Ni upgrade Ratio Mg dow ngrade Ratio 9

+ 8

7

6 13 5 O) c 4 5 § fi 2 3 S 2

+ 1

0 3 4 Mass Yield %

Figure 4.21. Ni Upgrade and Mg Downgrade Ratios versus Mass Yield

Figure 4.22 shows that the lower the concentrate mass yield, the higher the grade and the lower the Ni recovery.

Figure 4.22. Ni Grade/Recovery versus Cone. Mass Yield 61

The results indicate that the CVD can recover most of the fine particles of free gold, middlings containing gold grains, Ni bearing sulphides and also reject the Mg bearing minerals. For both Au and Ni, the grades decrease with increasing the mass pull but the recoveries increase.

4.2.2.3.3. Size Partition

The size partition curves were obtained based on the results of all pilot scale tests. The feed mass flow rate of each test was calculated based on measured mass flow rate of concentrate and tailing products. These data and the size analysis results were then used for calculating the weight percentages and the mass flow rates for each size fraction (i) of concentrate, tailing and feed (Appendix C.3). The partition number (PN) for each size fraction was determined as following:

PN; = 100 * (concentrate mass flow rate i /feed mass flow rate 0

Figure 4.23 shows the size separation partition curves for tests 3 and 7. For size enhancing the individual test results showed a good trend that CVD can recover the coarse middling particles with a cut size of about 250 micron. For test 7, the individual partition numbers for size fractions of +297, -297+210, and -210+149 pm are 81.1, 49.3 and 27.6, respectively. The results of size/assay analysis for all CVD tests are presented in Appendix C. The metal recovered was primarily in the coarse size fractions, which would likely occur as middling particles (sulfides + silicates). 62

90

80 •A- - - Test 3 - Test 7 m A ^ 70 i ;

JQ 60 Mil E 3 Z 50 c o S 40 MM a

°- 30

20 ! il 10-1 i! j I ! ! I! I 10 100 1000 Size (um)

Figure 4.23. Size Separation Partition Curves

4.2.2.3.4. Density Fractionation

The density fractionation tests on each size fraction of feed and concentrate of test 7 at densities of 2.7, 3.1, 3.5, 3.9 and 4.3 g cm"3 were performed using a Magstream. The density fractions were weighed and assayed. These data and the size analysis results were then used for calculating the weight percentages and the mass flow rates for each size fraction of concentrate (Appendix D.2). The same as feed mass flow rates, the concentrate mass flow rates for all density fractions of each size fraction (i) were also calculated based on concentrate mass flow rates of each size fraction and the weight percentage of each density fraction (Mass flow rate of each density fraction = Weight % * Mass flow rate of each size fraction). The 63 partition number for each density fraction was determined by using the following formula:

PN i = 100 * (concentrate mass flow rate j /feed mass flow rate >)

The tables of results are shown in Appendix D. The partition coefficient curves for densities were examined and presented in Figures 4.24 (+297 pm size fraction) and 4.25 (-297+149 pm size fraction). Figures 4.26 to 4.31 present the element recoveries for each density fraction.

20

10 -

0 J , , , , 2.25 2.75 3.25 3.75 4.25 4.75 Density (g/cm3)

Figure 4.24. Partition Coefficient Curve for +297 pm Size Fraction 100 -r

90 -

80

$ 70 n E 60 3 Z 50 • .2 40 1 30 • a Q- 20 ••

10

0 • 2.25 2.75 3.25 3.75 4.25 4.75 Density (g/cm3)

Figure 4.25. Partition Coefficient Curve for -297+149 um Size Fraction

Figure 4.26. Ni Recovery versus Density for +297 um Size Fraction Figure 4.27. Ni Recovery versus Density for -297+149 pm Fraction

Figure 4.28. Mg Recovery versus Density for-149+105 um Size Fraction 66

2.25 2.75 3.25 3.75 4.25 4.75 Density (g/cm3)

Figure 4.29. Mg Recovery versus Density for-105+74 um Size Fraction

The curves for density of individual size fraction showed that the heavy and coarser size fractions have sufficiently higher partition numbers than the lighter fine material. The Ni recovery is higher in the coarser and heavier fractions. The achieved results for Mg showed that the finer particle size fractions have the same partition numbers (less than 10%). The CVD separation performance revealed a good recovery for middlings (sulphides + silicates) and a good rejection of magnesium bearing minerals.

Figures 4.30 and 4.31 present sulfur and iron recoveries for densities of +297 pm size fraction. The iron and sulfur recovery in higher density of +297 pm fraction size have the maximum amount. It means that CVD has recovered the sulfides efficiently. 20 -1

10 0 -I , , , : , 2.25 2.75 3.25 3.75 4.25 4.75 Density (g/cm3)

Figure 4.30. S Recovery versus Density for +297 um Size Fraction

100

10 -I 0 J , , , , 2.25 2.75 3.25 3.75 4.25 4.75 Density (g/crr>3)

Figure 4.31. Fe Recovery versus Density for +297 um Size Fraction 68

4.2.2.4. Flotation Test Results and Proposed Flowsheet

The representative portion of CVD concentrate product (Pgn = 160 pm) obtained from

Tests 7 of SC-4 sample was ground to 80% less than 100 pm (based on data from existing plant). The flotation test was conducted on the ground sample at pH 10 and

addition of PAX (125 g/t) and MIBC. Soda ash (Na2C03) was used as a pH regulator.

The flotation test conditions and results are shown in Appendix E. The aim of the test

was to float the nickel sulphides, which tends to maximize the recovery of the nickel.

The metallurgical balance is shown in Table 4.13.

Table 4.13. Flotation Test Results

weight Grade Distribution % Material Ni Fe Mg S Au % Ni Fe Mg S Au % % % % (g/t) Concentrate 35.1 1.32 50.4 2.7 29.7 3.97 64.1 37.6 26.1 38.9 98.2 Tailing 64.9 0.40 45.2 4.1 25.2 0.04 35.9 62.4 73.9 61.1 1.8 Feed(CVD Con) 100 0.72 47.0 3.6 26.8 1.42 100 100 100 100 100

The nickel recovery of the flotation test is about 64%, which can have an impact

on overall nickel and gold recoveries of the beneficiation plant. Based on the existing

plant results (80% nickel recovery), CVD Test 7 (16.1% recovery from SC-4 tailings)

and followed by flotation (nickel recovery of 64.1%), the overall nickel recovery will be

increased by about 1%. In case of gold the overall gold recovery will be increased by

6%. The proposed flowsheet is shown in Figure 4.32. 69

AS-4 Feed Roughers ,r fc Regrind Scavengers w Rougher-Scavenger Tails Rougher Con

Rougher-Cleaner Tails SC-4 Rghr-Clnrs Scav-Clnrs

Scavenger-Cleaner Tails Xu/Ni Con Cu/Ni Tails Smelter Ni Con Cu/Ni Sep. •0 • Knelson Cu-Cleaners Tails CVD Cu-Clnrs

T Smelter Cu Con Tailings

Figure 4.32. Proposed Flowsheet, INCO Thompson 70

4.2.3. Conclusion and Recommendation

Two samples of flotation tailings referred to as Scavenger-Cleaner (SC-4) and Scavenger (AS-4) from the Birchtree ore processed at the Thompson beneficiation plant, was studied. The test program was directed into three stages including sample characterization, batch gravity and pilot scale CVD testing followed by product characterization and density fractionation of sized fractions of feed and concentrate.

The data from sample characterization and batch tests on samples revealed that both samples have potential for middling recovery. The Pso for SC-4 sample is 94 um and for AS-4 is 140 um. A batch 3 inch Knelson was used to obtain a preliminary assessment of nickel recovery. For the SC-4 sample, the results were similar to the target results based on the size-assay analysis. For the AS-4 sample, while the Ni grade was similar to the target value, the weight and Ni distributions were lower. The Mg grade is decreased dramatically in the concentrate products. For SC-4, only 3.3% of Mg reported to the concentrate product.

Seven pilot scale tests were performed on the SC-4 sample using the CVD6 separator. Operating conditions including pinch valve closed time and bowl speed were varied producing a range of results. The pinch open time was maintained at 0.03 second and the fluidization water flow rate maintained at 8 gallon per minute for all tests. Products were subjected to size/assay analysis to assess separation. The results showed the possibility of producing a concentrate with a reasonable low mass yield, low magnesium and high nickel upgrade ratio.

Pilot scale testing indicated that the CVD was effective at recovering nickel sulphides from flotation tailings. The results indicated that the Ni recovered was primarily in the coarse size fractions, which would likely contain middling particles (sulfide + silicates) i.e. size enhanced density separation. The best result (Test 3) was 6.7% concentrate mass yield with 0.85% Ni grade and 13% Ni recovery. The Mg distribution in concentrate was just 1.7%. This is a good indication that the centrifugal 71 concentrator can also remove the magnesium bearing minerals such as talc. In case of gold remarkable results were produced and the metallurgical balances showed the gold recovery of 67% for this test.

Comparison of nickel, sulphur, and gold distribution indicates that free gold is being recovered. According to the results attained, the CVD machine can upgrade the flotation tailings of the Birchtree ore processed at the Thompson beneficiation plant with a reasonable recovery. It seems that the CVD recovers not only the coarser grains of nickel bearing sulphides but also captures the ultrafine free gold particles.

Size partition curves were obtained for all tests, which demonstrated that the CVD

operates as a size classifier as well as a density separator (size enhanced density

separation). For Test 7, the cut size was about 250 microns and the individual partition

numbers for size fractions of +297, -297+210, and -210+149 pm are 81.06, 49.33 and

27.64, respectively. The curve for density of individual size fraction did not show a

systematic trend for test 7, at the same time the partition numbers for the heavier portion

of coarser size classes (+50, -50+100 mesh) are high.

The results showed that the process option to recover the middling particles

containing nickel sulphides and fine gold particles is application of the CVD for the SC-4

sample. It seems that the CVD recovers not only the coarser grains of Ni bearing

sulphides but also captures the ultrafine free gold particles simultaneously.

The results also indicated a product with very low content in Mg is achievable.

Recycling the concentrate product of CVD to the grinding mill may improve the overall

recovery. The separation efficiency performance is being affected by the feed

characterization.

Flotation test on ground CVD concentrate of SC-4 sample showed that with

recycling the CVD concentrate to the grinding and flotation circuit, the overall nickel and

gold recoveries will be increased by about 1% and 6%, respectively. 72

Modeling of grinding, flotation and CVD circuits to assess potential improvements in overall metal recovery and assess the potential for increasing grind size to reduce losses in ultrafine fraction, reduce energy requirements and increase throughput is recommended.

Based on the low mass stream of the scavenger-cleaner tailings at the Thompson beneficiation plant (for Birchtree ore), installation of a Knelson CVD6 is recommended to perform plant-trial test. It can be located inside the grinding and flotation circuit to minimize overgrinding of materials. 73

CHAPTER 5 Conclusions and Recommendations

5.1. Conclusions

A study was conducted to evaluate the Knelson Continuous Variable Discharge (CVD) concentrator as a scavenger for coarse middling particles from flotation tailings. The tailing samples were from Eskay Creek and INCO Thompson beneficiation plants. The goal was to produce a product suitable for recycling to grinding to improve the degree of liberation and aid subsequent recovery in flotation.

• Based on the characterization and batch Knelson tests, there appeared to be an opportunity to recover a recyclable product from the flotation tailings.

• Pilot scale testing indicated that the CVD was effective at recovering gold and sulphides from flotation tailings. The results indicate that the gold and sulphides recovered were primarily in the coarse size fractions, which would likely contain middling particles (sulfide + silicates) i.e. size enhanced density separation.

• For Eskay Creek sample (LGT) the best result (test 3) was 3.3 % mass yield with 12.5 ppm Au grade and 28% Au recovery. For this test the Au upgrade ratio was 8.7.

• For INCO Thompson sample (SC-4) the best result (test 3) was 6.7% concentrate mass yield with 0.85 % Ni grade and 13% Ni recovery. The Mg distribution in concentrate was just 1.7%.

• Comparison of sulphur, silver and gold distribution indicates that free gold is being recovered. According to the results attained, the CVD machine can upgrade the flotation tailings of the Eskay Creek gold processing plant with a reasonable recovery. It seems that the CVD recovers not only the coarser grains of gold associated sulphides but also captures the ultrafine free gold particles. 74

Size partition curves were obtained for all tests, which demonstrated that the CVD operates as a size classifier as well as a density separator (size enhanced density separation). For tests 3 and 6 (Eskay) and test 7 (INCO), the cut size was about 300 and 250 microns, respectively.

Flotation test on ground CVD concentrate of LGT and SC-4 samples showed that with recycling the CVD concentrate to the grinding and flotation circuit, the overall nickel and gold recoveries will be improved. .75

5.2. Recommendations

• Modeling of grinding, flotation and CVD circuits to assess potential improvements in overall metal recovery and assess the potential for increasing grind size to reduce losses in ultrafine fraction, reduce energy requirements and increase throughput is recommended.

• Based on the low mass stream of tailings at the Eskay Creek beneficiation plant, installation of Knelson CVD6 is recommended to perform plant-trial test. It can be located inside the grinding and flotation circuit to minimize overgrinding of materials. Even at the Eskay Creek the Knelson batch machine has been located and fed from cyclone underflow discharge, The CVD can upgrade the flotation tailings at a low mass yield.

• Based on the low mass stream of the scavenger-cleaner tailings (SC-4) at the Thompson beneficiation plant (for Birchtree ore), installation of a Knelson CVD6 is recommended to perform plant-trial test. 76

REFERENCES

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Awmack J., B. Kuljit, R. Buhler, F. Jesse, 2004, "The effectiveness of the Knelson CVD to process Flotation Tailings from the Eskay Creek Mill", Mine 338 Course, Mining Engineering Department, UBC, pp. 35-47.

Burt, R.O., 1984, "Gravity Concentration - from Bench Scale to Plant", Annual Meeting of Canadian Mineral Processors, Ottawa.

Burt, R.O., G. Korinek, SR. Young, C. Deveau, 1995, "Ultrafine Tantalum Recovery Strategies", Minerals Engineering, Vol. 8, pp 859-870.

Burt, R.O., 1999 "The Role of Gravity Concentration in Modern Processing Plants", Minerals Engineering, Vol. 12, No. 11, pp 1291-1330.

Burt, R.O., 1998,"A Review of Gravity Concentration Techniques for Processing Fines, Production and Processing of Fine Particles", CIM, pp 375-385.

Byron R, K. Roberts, 2004, "Flotation Improvements in the Luzenac Penhorwood Talc Concentrator", Proceedings 36th Annual Meeting of the Canadian Mineral Processors, Ottawa, Canada, pp. 177-188.

Chan BS. K., R.H. Mozley and G.J.C. Childs, 1991, "Extended trials with the high tonnage Multi-Gravity Separator", Minerals Engineering, Vol. 4, No. 3/4, pp. 489- 496. 77

Drapack J., 2004, "Identification of a Birchtree Ore Mill Stream for Application of the Knelson CVD Concentrator for Nickel and Precious Metals Recovery", Internal

Report for Dr. Klein, Department of Mining Engineering, University of Britrish

Columbia.

Guerney P. J., A. R. Laplante S. O'Leary, 2003, "Gravity recoverable Gold and the

Mineral Liberation Analyzer", Proceedings 35lh Annual Meeting of the Canadian

Mineral Processors, pp.401-416.

Hilliard T., 2003, "TANCO - Endurance through adaptability", CIM Bulletin, Vol. 96 No. 1070, pp.43-46.

Honaker R. Q, B. C. Paul, D. Wang and M. Huang, 1995, "The application of Centrifugal Washing for Fine Coal Cleaning", Minerals & Metallurgical Processing, Vol. 12, pp. 80-84.

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Butterworth - Heineman Publications, Oxford, England, pp. 16-18.

Klein B , 2002, "Processing of Precious Metal Ores (Course Notes)", Department of

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New Continuous Centrifugal Concentrator", Presentation at BC and Yukon CMP

Annual Meeting. 78

Laplante, AR., J. Ling and J. Xiao, 1998, "Difficult Gravity Separation an Update",

30th Annual Meeting of Canadian Mineral Processors, Ottawa, Paper 35, pp. 619-637.

Lynch A.J., N.W. Johnson, E.V. Manlapig, C.G. Thorne, 1981, "Mineral and Coal Flotation Circuits, their Simulation and Control (Development in Mineral Processing)" Elsevier Scientific Publishing Company, Amsterdam, The Netherlands, pp. 33-56.

Mcleavy M., B. Klein and I. Grewal, 2001, "Knelson Continuous Variable Discharge Concentrator, Analysis of Operating Variables", International Heavy Minerals Conference, Fremantle, Australia, pp. 119-125.

Mohanty M. K. and R. Q. Honaker., 1999, "Evaluation of the Altair Centrifugal Jig for Fine Coal Processing", Coal Preparation, Vol.20, pp. 85-106.

Mohanty M. K and R. Q. Honaker., 1998, "Evaluation of the Altair Centrifugal Jig for Fine Particle Separation", SME Annual Meeting, Orlando Florida, pp. 1-11.

Napier-Munn T.J., S. Morrell, R.D. Morrison, T. Kojovic, 1999, "Mineral Comminution Circuits, Their Operation and Optimization", Julius Kruttschnitt Mineral Research Centre, The University of Queensland, Australia, , pp. 349-352.

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Wyslouzil, H. E., 1990, "Evaluation of the Kelsey Centrifugal Jig at Rio Kemptville

Tin", 22nd Annual Meeting of the Canadian Mineral Processors, Ottawa, Paper No 23, pp. 461-472. 80

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Report", Canadian Milling Practice, CIM Special Volume 49, pp. 101-103.

Falcon Concentrators http://www.concentrators.net/

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IGC Corporation, February 1991, "Operating Instructions for the Magstream Models

50 and 100 Laboratory Separators". Appendices

Appendix I - Eskay Creek 82

Appendix IA Sample Characterization

A.l Particle Size Distributions

Table A.1.1. LGT Size Analysis

Size Fraction Weight % cummul. Mesh um 9 % Passing +50 +297 2.5 0.1 99.9 -50+70 -297+210 20.4 1.0 98.9 -70+100 -210+149 177.0 8.2 90.7 -100+140 -149+105 432.4 20.1 70.6 -140+200 -105+74 401.8 18.7 51.9 -200+270 -74+53 201.5 9.4 42.5 -270+400 -53+37 90.8 4.2 38.3 -400 -37 821.4 38.3 Tot. 2147.7 100.0

v . - -,

90

80

0> 70

| 60 CO ^ 50

E 40

5 30

20

10

0 10 100 1000 Size (um)

Figure A.1.1. Particle Size Distribution, LGT 83

Table A.1.2. HGT Size Analysis

Size Fraction Weight % cummul. Mesh um g % Passing +50 +297 12.0 0.6 99.4 -50+70 -297+210 13.2 . 0.7 98.7 -70+100 -210+149 76.8 4.0 94.7 -100+140 -149+105 235.5 12.3 82.4 -140+200 -105+74 242.9 12.7 69.7 -200+270 -74+53 229.0 12.0 57.7 -270+400 -53+37 152.2 7.9 49.8 -400 -37 953.6 49.8 Tot. 1915.2 100.0

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20 x

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10 100 1000 Size (Lim)

Figure C.6.1. Size Separation Partition Curves, LGT

Au Recovery versus Size 102

10 100 1000 Size (um)

Figure C.6.3. S Recovery versus Size, LGT

x

S Recovery %

Figure C.6.4. Au Recovery versus S Recovery, LGT Appendix ID Mineralogical Analysis fo3-l

Concentrate +100 Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 70 7 2.985 208.95 45.73 21-40 0.3 25 7.5 3.655 91.38 20.00 41-60 0.5 10 5 4.325 43.25 9.47 61-80 0.7 0 0 4.995 0.00 0.00 81-100 0.9 20 18 5.665 113.30 24.80 Total 125 37.5 456.88 Total 1000 Grains

Class: Cone.+100 Texture: Complex Grain Shape: Sub-Rounded Grain Size: >149 pm Mineral Associations: Primarily associated with the light colored gangue mineral Mineral Surface: Because the sulfides exist as fine inclusions, it appears as though there in insufficient surface area, indicating lowest sulfide recovery. Impurities or Inclusions: Sulfide exists as fine inclusions within the gangue particles

Cone +100 Class Middlings Distribution

0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of liberation \ol-Z

Concentrate -1 00 Weight Wt% Class Average # Of Sulphides Weight per grain 223.88 24.08 0-20 0.1 75 7.5 2.985 21^0 0.3 45 13.5 3.655 164.48 17.69 41-60 0.5 35 17.5 4.325 151.38 16.28 49.95 5.37 61-80 0.7 10 7 4.995 81-100 0.9 60 54 5.665 339.90 36.57 929.58 Total 225 99.5 Total 2500 Grains

Class: Cone.-100 Texture: Fairly complex. Some liberated particles Grain Shape: Sub-Rounded Grain Size: > 74 pm, < 149 pm Mineral Associations: Primarily associated with the light colored gangue mineral Mineral Surface: Because the sulfides exist as fine inclusions, it appears as though there in insufficient surface area, indicating low sulfide recovery. Impurities or Inclusions: Sulfide exists as fine inclusions within the gangue particles

Cone -100 Class Middlings Distribution

40.00 35.00 30.00 25.00 20.00 15.00 10.00 5.00 0.00 0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 49.09 Concentrate -200 Weight Class Average # Of Sulphides Weight per grain Wt% 0-20 0.1 100 10 2.985 298.50 15.18 21-40 0.3 60 18 3.655 219.30 11.15 41-60 0.5 50 25 4.325 216.25 10.99 61-80 0.7 20 14 4.995 99.90 5.08 81-100 0.9 200 180 5.665 1133.00 57.60 Total 430 247 1966.95 Total 2000 Grains

Class: Cone.-200 Texture: Primarily simple. Few complex particles, mainly liberated particles. Grain Shape: Both Sub-Angular and Sub-Rounded Grain Size: > 37 pm, < 74 pm Mineral Associations: There are more particles which are fully liberated, hence minimal association Mineral Surface: Because most particles are fully liberated, recovery is higher Impurities or Inclusions: Primarily none

Cone -200 Class Middlings Distribution

70.00 60.00 50.00 40.00 30.00 20.00 10.00 0.00 0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 75.90 Concentrate -400 Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 60 6 2.985 179.10 1.95 21-40 0.3 20 6 3.655 73.10 0.80 41-60 0.5 80 40 4.325 346.00 3.77 61-80 0.7 120 84 4.995 599.40 6.53 81-100 0.9 1410 1269 5.665 7987.65 86.96 Total 1690 1405 9185.25 Total 5000 Grains

Class: Cone.-400 Texture: Fully Liberated, primarily simple. Grain Shape: Both Sub-Angular and Sub-Rounded Grain Size: < 37 |Jm Mineral Associations: None Mineral Surface: Fully liberated, so highest recovery Impurities or Inclusions: None

Cone -400 Class Middlings Distribution

100.00 —i

80.00 pm 60.00 m i I m i M i — .— i—n m i 0-20 21-40 41-60 61-80 81-100

Class Middlings

Degree of Liberation = 90.32 '03-5"

Feed +100 Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 110 11 2.985 328.35 65.29 21-40 0.3 30 9 3.655 109.65 21.80 41-60 0.5 15 7.5 4.325 64.88 12.90 61-80 0.7 0 0 4.995 0.00 0.00 81-100 0.9 0 0 5.665 0.00 0.00 Total 155 27.5 502.88 Total 1000 Grains

Class: Feed+100 Texture: Complex Grain Shape: Sub-Angular Grain Size: >149 \im Mineral Associations: Associated with both light and dark colored gangue minerals Mineral Surface: Low sulfide content. Because the sulfides exist as fine inclusions, it appears as though there in insufficient surface area for flotation of the sulfides Impurities or Inclusions: Sulfide exists as fine inclusions within the gangue particles.

Feed +100 Class Middlings Distribution

70.00 60.00 50.00 gS 40.00 5 30.00 20.00 10.00 0.00 0-20 21-40 41-60 61-80 81-100

I Class Middlings

Degree of Liberation = 0 Feed -100 Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 90 9 2.985 268.65 49.00 21-40 0.3 25 7.5 3.655 91.38 16.67 41-60 0.5 5 2.5 4.325 21.63 3.94 61-80 0.7 5 3.5 4.995 24.98 4.56 81-100 0.9 25 22.5 5.665 141.63 25.83 Total 150 45 548.25 Total 2500 Grains

Class: Feed-100 Texture: Lower degree of complexity. Grain Shape: Both Sub-Angular and Sub-Rounded Grain Size: > 74 |jm, < 149 |Jm Mineral Associations: Associated with both light and dark colored gangue minerals Mineral Surface: Low sulfide content. Because the sulfides exist as fine inclusions, it appears as though there in insufficient surface area for flotation of the sulfides Impurities or Inclusions: Sulfide exists as fineinclusion s within the gangue particles

Feed -100 Class Middlings Distribution

21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 50.0 Feed -200 Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 230 23 2.985 686.55 47.72 21-40 0.3 50 15 3.655 182.75 12.70 41-60 0.5 40 20 4.325 173.00 12.02 61-80 0.7 0 0 4.995 0.00 0.00 81-100 0.9 70 63 5.665 396.55 27.56 Total 390 121 1438.85 Total 2000 Grains

Class: Feed-200 Texture: Primarily simple. Some fully liberated particles Grain Shape: Both Sub-Angular and Sub-Rounded Grain Size: > 37 pm, < 74 pm Mineral Associations: Mainly associated with light gangue mineral. There are more particles which are fully liberated Mineral Surface: Because more particles are fully liberated, higher sulfide content Impurities or Inclusions: Sulfide exists as both fineinclusion s within the gangue particles and as liberated particles

Feed -200 Class Middlings Distribution

60.00

0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 52.07 Feed -400

Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 140 14 2.985 417.90 13.78 21-40 0.3 40 12 3.655 146.20 4.82 41-60 0.5 40 20 4.325 173.00 5.70 61-80 0.7 40 28 4.995 199.80 6.59 81-100 0.9 370 333 5.665 2096.05 69.11 Total 630 407 3032.95 Total 5000 Grains

Class: Feed-400 Texture: Fully Liberated Grain Shape: Sub-Rounded Grain Size: < 37 pm Mineral Associations: None Mineral Surface: Fully liberated, indicating highest sulfide content of the feed Impurities or Inclusions: None

Feed -400 Class Middlings Distribution

80.00

60.00

40.00

20.00

0.00 0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 81.82 Tails +10C Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 100 10 2.985 298.50 80.33 21-40 0.3 20 6 3.655 73.10 19.67 41-60 0.5 0 0 4.325 0.00 0.00 61-80 0.7 0 0 4.995 0.00 0.00 81-100 0.9 0 0 5.665 0.00 0.00 Total 120 16 371.60 Total 1500 Grains

Class: Tails+100 Texture: Complex Grain Shape: Sub-Rounded Grain Size: >149 pm Mineral Associations: Associated with both light and dark colored gangue minerals Mineral Surface: Little to no sulfide content Impurities or Inclusions: Sulfide exists as fine inclusions within the gangue particles

Tails +100 Class Middlings Distribution

100.00

80.00

3= 60.00

§ 40.00

20.00

0.00 0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 0 Tails -100 Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 95 9.5 2.985 283.58 44.68 21-40 0.3 20 6 3.655 73.10 11.52 41-60 0.5 25 12.5 4.325 108.13 17.03 61-80 0.7 0 0 4.995 0.00 0.00 81-100 0.9 30 27 5.665 169.95 26.77 Total 170 55 634.75 Total 1000 Grains

Class: Tails-100 Texture: Lower degree of complexity. Grain Shape: Sub-Rounded Grain Size: > 74 pm, < 149 pm Mineral Associations: Associated with both light and dark colored gangue minerals Mineral Surface: Low sulfide content Impurities or Inclusions: Sulfide exists as fine inclusions within the gangue particles

Tails -100 Class Middlings Distribution

50.00

40.00

3? 30.00

§ 20.00

10.00 0.00 0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 54.27 Tails -200 Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 150 15 2.985 447.75 38.26 21-40 0.3 50 15 3.655 182.75 15.62 41-60 0.5 10 5 4.325 43.25 3.70 61-80 0.7 20 14 4.995 99.90 8.54 81-100 0.9 70 63 5.665 396.55 33.89 Total 300 112 1170.20 Total 2000 Grains I

I

( Class: Tails-200 Texture: Primarily simple. Some fully liberated particles 1 Grain Shape: Both Sub-Angular and Sub-Rounded i Grain Size: > 37 pm, < 74 pm Mineral Associations: Mainly associated with light colored gangue mineral. There are more particles which are fully liberated, indicating the CVD was unable to recover these sulfide particles. Mineral Surface: Because more particles are fully liberated, higher sulfide content Impurities or Inclusions: Primarily none due to simple structure

Tails -200 Class Middlings Distribution

50.00 T-

0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 72.87 Tails -400 Class Average # Of Sulphides Weight per grain Weight Wt% 0-20 0.1 80 8 2.985 238.80 8.44 21-40 0.3 50 15 3.655 182.75 6.46 41-60 0.5 80 40 4.325 346.00 12.23 61-80 0.7 50 35 4.995 249.75 8.82 81-100 0.9 320 288 5.665 1812.80 64.05 Total 580 386 2830.10 Total 5000 Grains

Class: Tails-400 Texture: Fully Liberated Grain Shape: Both Sub-Angular and Sub-Rounded Grain Size: < 37 pm Mineral Associations: None Mineral Surface: Fully liberated. Indicating the CVD was unable to recover these sulfide particles, and highest sulfide content of the tails. Impurities or Inclusions: None

Tails -400 Class Middlings Distribution

70.00 60.00 50.00 3? 40.00 S 30.00 20.00 10.00 0.00 0-20 21-40 41-60 61-80 81-100 Class Middlings

Degree of Liberation = 90.32 104

Appendix IE Flotation Test

Flotation Test on CVD Concentrate

Sample: CVD Concentrate (CVD Pilot Test 6) Sample weight: 500 g

Particle Size: P8o = 180 (im

For flotation test the sample was ground to Pso = 75 pm.

Flotation Procedures and Conditions:

pH: Natural Conditioning time: 5 minutes Collector: PAX (100 g/t) Frother: MIBC Rougher flotation time: 5 minutes

Table 1. Metallurgical Balance, LGT

Weight Grade Distribution % Material % Au (ppm) Ag (ppm) Fe% S % Au Ag Fe S Concentrate 33.4 9.26 141.0 3.85 2.3 81.0 72.9 37.6 74.3 Tailings 66.6 1.09 26.3 3.2 0.4 19.0 27.1 62.4 25.7 Feed (CVD Con) 100 3.82 64.6 3.42 1.03 100 100 100 100 Appendix II - INCO Thompson Appendix HA Sample Characterization

A.l Particle Size Distributions

Table A. 1.1. SC-4 Size Analysis

Size Fraction Weight % cummul. Mesh um 9 % Passing +50 +297 127.1 3.0 97.0 -50+70 -297+210 113.2 2.7 94.3 -70+100 -210+149 192.3 4.5 89.8 -100+140 -149+105 289.4 6.8 83.0 -140+200 -105+74 460.6 10.8 72.2 -200+270 -74+53 755.9 17.8 54.4 -270+400 -53+37 488.5 11.5 42.9 -400 -37 1824.6 42.9 Tot. 4251.7 100.0

. .4

80 4 /

20

g ^ I I I I I I I I | I I I I I I I I | 10 100 1000 Particle Size(um)

Figure A.1.1. Particle Size Distribution, SC-4 107

Table A.1.2. AS-4 Size Analysis

Size Fraction Weight % cummul. Mesh um g % Passing +50 +297 86.5 1.5 98.5 -50+70 -297+210 233.2 3.9 94.6 -70+100 -210+149 676.0 11.4 83.2 -100+140 -149+105 994.6 16.8 66.4 -140+200 -105+74 1004.6 17.0 49.5 -200+270 -74+53 1117.0 18.9 30.6 -270+400 -53+37 532.1 9.0 21.6 -400 -37 1281.5 21.6 Tot. 5925.4 100.0

100

! I ! ! -:* I i : I 1 i i ; .•: ] ! 80 i I ! 1 i !: i 1 / I ; i I

: ill! i i ji.il.*.- I i | !!;!•' ! | ! • 1 i I ; i : /! i!! 4 ! ! i 11 E I

i | i III! ' I ! i ! ! MM 1 | ! | i 20 • i ! I! 1; ! ! i i i \ i i ' i i \ i ; ! i 1 0 10 100 1000 Particle Size (um)

Figure A.1.2. Particle Size Distribution, AS-4 CO CO

CO 05

CO

CM

CO LO LO CO

CO CO

3 of 'I

ID CO LO LO H CO

CM CN CM CT> < CU s co CM CO H CO CO A.3 Magnetic Particles Distribution

Table A.3.1. Magnetic Particles Distribution, SC-4

Size Non Magnetics Magnetics Fraction magnetics 9 % (um) 9 +297 2.1 76.9 2.6 -297+210 12.6 56.1 18.3 -210+149 32.2 111.0 22.5 -149+105 48.2 184.2 20.7 -105+74 75.8 186.2 28.9 -74+53 50.7 115.5 30.5 -53+37 50.8 101.8 33.3 -37 27.7 177.0 13.5 Tot. 299.9 1008.6 22.9

50

40 ]

I JZL +297 -297+210 -210+149 -149+105 -105 +74 -74 +53 -53 +37 -37 Size Class (um)

Figure A.3.1. Magnetic Particles Distribution for each size fraction of SC-4 Table A.3.2. Magnetic Particles Distribution, AS-4

Size Non Magnetics Magnetics Fraction magnetics g % (pm) g +297 0.9 35.2 2.5 -297+210 18.2 152.5 10.6 -210+149 47.2 273.5 14.7 -149+105 47.3 137.7 25.6 -105+74 60.3 153.2 28.2 -74+53 79.7 198.3 28.7 -53+37 64.5 113.9 36.1 -37 53.9 122.8 30.5 . Tot. 371.9 1187.1 23.9

10 4

+297 -297+210 -210+149 -149+105 -105+74 -74 +53 -53 +37

Size Class (um)

Figure A.3.2. Magnetic Particles Distribution for each size fraction of AS-4 Tt CN oq Tf m rn d d d CN •n Tf SO 21. 2 Ma g p o c op rn oo o o ca in CN p r-- O CN rn CN 00 CN od CN Tf r~ z s 60 CN Tf oo p Tf ca Tf d d rn rn SO CN p 00 o o a oo in rn 00 oo •n CN Tf s o ca r~ od CN CN Tt SO m r- z s oo CN ca m sq rn SO Tf oo p in d d CN rn •n CN CN p

ution % 3 o CJ s o B OO oq in O o Tf o Tf od o ca TT t~ Tf rn CN SO SO Tf Distri b Z S 60 m OO rn OO Os sq CN CN ca d d d CN •n rn SO CN p o C/3 o 7 c s jp in Tf oq o ca rn rn in so rn CN TT rn r-" Os' od u Z 2 m 00 CN »> TT Os sq rn CN sq ca in d d d rn rn so CN p "a o z s Os oq c« oc cSaP 00 Os sq m Tf os CU d od CN CN rn TT Tf rt Z S m r~

« 00 O o Tf CN r- Os in Tf CN oo Tf in rn CN sq ca Tt od OS in od CN od CN CN CN m m .a 2 CN CN CN CZ) u. CN SO OS m m Os SO t-; OO OS Os cn o ca t-; O oo SO

a •n OS oo m m Tf OS m CU SP •n in SO in Tf m m Tf CD o o p o o O o o cu d d d d d d d d CU ide % 3 d CJ s So CN m so oo in O TT TT Tf m CN CN CN Tf rn « o ca o O O o O o o o O d d d d d z s d d d d 00 oo Tf in r- oo Tf Tf oo OO rn OS ca r-; Tf Tf rn Tf Tf TT m' rn rn Tf rn

— m so O in r~ o m oo O rt SO TT oo o Tf oo m o Tf Tf 3 ZS Tr' CN CN rn TT Tf rn cct H OS m CS TJ" m Tf oo CN o CN SO •n 00 Tf so sri SO in Tf in •n in sq d d d d d s d d d d z in •n in B 60 m m oo m o m OS oo r- oo o Tf Tj- m rn Tf o ca m CN CN m SO d d d d d d d d z s d oo OS CN m m CN oo 60 o Tf O Tt Tf OO oq i—< in in ca d d rn rn CN -*—» rf s p od m o Weig l o o o OS in rn sq oq TT CN od CN CN rn •n m NonMag %

o Os iv, Tf O Tf rn r~ CN 6 OS + + + + + o Os + m m

Siz e o f- 3. + OS Tf Tf CN o Fractio n CN m OO CN Os CN d d rn TT' od rn 28. 0 Ma g O 1> o c op Os Os oq p o ca p 00 sq p o m" CN TT od CN CN Z £ - CN

60 00 m 00 p CN p ca od £ d d Tr" sd TT Tt CN iri p 00 CN o £ u-i oo o c op rn sq Os p Os O Z £ rn oc CN CN iri SO CN

00 sq Os rn sq oo p ca m r-- d d rn w~i u-i TT od £ CN p

ution % 3 o u c op r- p o ca ON o 00 p oo oo o TI" CN oo OS (--* Tt CN

Distri b z £

00 oo p sq CN Os sq p ca od d d CN rn Tf SO rn sd £ CN o o CN rn O o c op Tf p Os p CN o * i/-i r~* d CN CN Z £ CN rn CN OS CN

00 m sq ST) oo CN sq sq rn p ca 00 d d rn l/-i r-" «-i £ CN CN CN p o Z o c op C-; p SO OO OS CN m O o ca sd CN od Z £ rn SO O* Os SO CN f~

SO m r- oo CN Tt CN r- 00 TT iri C3 OS sq r- CN OS •

00 CN Tf IT) 1/1 OS OS Os Os ca Tf oo l/-i iri d £ Os d r-^ d 6X1 t-H £ c OS SO 1/-S l/"» m OS m Os 2P Tt p rn o <« o rn m p Z £ CN TT rn CN rn CN rn m"

OO r~ SO m in i/-i m m m CN CN o o o CN O o CN ca o o d d d d o d d o £ d

ade % 3 d d

G r u -3- un m SO SO CN m O CN CN o o CN CN o O O o O o d d d d d d d d d

OS oo OS o oo Tl" . gp OS >/-> m SO £ CN Tf CN CN* CN rn Tf' Tf' rn

«-i TJ- CN TT CN SO Os Os CN o ca Tf O oo Tf Os rn TT CN oo Tt rn rn rn z £ d CN

Vi CN O T1- OS Tf TT in OS C^ TT m o OS 00 m CN CN CN m m CN CN CN £ d d d d d d d d d

z 00 o o CN «-> oo o ca I/I m Tt SO SO OS OS CN rn Z £ d d d d d d d d d

so oo CN •>*• SO Tl" so 00 Os in OS oo 00 t-^ ca d d CN* Tr" Tt" sd CN vi CN £ o d o T Os Tf O Tl- Weig l o SO OS TT m p p O O rn TT CN TT CN CN CN sd CN* rn oo SO NonMag %

V, o OS TT TT rn r- o m S + + + + Os + TT +

Siz e 3. OS o O r~ 1 + (N Os "? Fractio n •e *> s .2 w S 3 Of .0 Density Grade Distribution % 0 ? £ S? z S Class Ni% Stot.% Fe% Cu% Au (ppm) Stot. Fe M Cu Au

d<2.7 55.57 0.195 2.71 10.84 7.22 0.042 0.013 37.0 43.0 43.5 46.2 50.0 16.4 O C od 2.7

d>4.3 10.34 0.775 5.24 16.76 11.32 0.083 0.263 27.4 15.5 12.5 13.5 18.3 60.3

1 1 000 1 1 00 0 1 00

0.293 3.50 13.86 8.69 0.047 0.045 100.0 100.0 100.0 100.0 0

d<2.7 40.72 0.157 1.45 6.46 7.77 0.031 0.010 18.5 13.1 21.1 28.5 29.7 13.6

2.7"

d>4.3 25.99 0.757 11.68 23.87 13.33 0.054 0.069 56.8 49.8 31.2 33.2 60.4

1 1 00 0 1 1 00 0 1 000 1 00 0 1 00 — 0 0.346 4.51 12.46 0.043 0.030 100.0 100.0 0 Ov d<2.7 ' 40.21 0.201 3.17 10.38 0.035 0.017 27.3 20.8 24.0 35.1 34.9 26.8 V rn V 19.61 0.269 3.98 14.85 11.09 0.045 0.027 17.8 12.7 16.7 21.3 21.9 21.1 m 0 3.1 0 »n <> rs vo rs -149+105 6.81 3.5

d>4.3 30.07 0.445 12.24 29.64 10.43 0.041 0.030 45.3 60.1 51.2 30.8 30.3 36.2

1 1 00 0 1 1 00 0 1 00 0 1 000 1 00

0.295 6.13 17.41 10.20 0.041 0.025 100.0 100.0 0 .c

Oil o Grade Distribution % Size SD <*" S? v Density OX) a Z Class(um) Class Ni% Stot.% Fe% Cu% Au (ppm) Stot. Fe BC Cu Au rn d<2.7 41.49 0.155 2.42 7.92 9.55 0.022 0.011 27.0 15.3 44.1 35.5 25.7 - p od 2.7

d>4.3 40.50 0.320 14.43 37.54 7.18 0.026 0.020 54.4 76.0 70.7 32.3 40.2 46.6

1 1 00 0 1 00 0 1 00 0 1 1 000 1 1 00

0.238 7.68 21.50 8.99 0.026 0.017 100.0 100.0 0

d<2.7 41.03 0.165 3.25 10.25 9.93 0.020 0.071 27.5 15.7 17.2 49.2 36.6 73.9 r-' oq 2.7

d>4.3 40.29 0.334 15.58 40.74 5.84 0.022 0.013 54.8 73.7 67.3 28.4 38.9 12.8

1 1 000 1 00 0 1 1 00

100.0 0.245 8.52 24.39 8.28 0.023 0.039 100.0 100.0 100.0 0

1 1 0 6 rn r~ 6

42.91 0.659 4.47 20.55 0.044 0.032 57.2 45.7 45.9 45.5 47.4 60.1

1 1 00 0 1 1 00

Tot.(calc) 100.0 0.434 6.09 20.78 9.86 0.036 0.032 100.0 100.0 100.0 100.0 0 <* 3 .c s •s §^ U 0- Mo Grade Distribution % Density u ? z

Class M Ni% Stot.% Fe% Mg% Au (ppm) Stot. Fe Cu Au

d<2.7 37.1 0.166 1.05 5.57 9.04 0.033 0.023 20.9 17.7 26.0 32.0 33.4 10.3 rn o- 2.7n rr' d CN m rn p rn r rr' O 3.9

d>4.3 23.01 0.684 5.11 12.99 13.28 0.043 0.032 53.4 53.3 37.6 29.1 26.5

1 1 00 0 1 1 00 0 1 1 00 0 1 00

100.0 0.295 2.206 7.943 10.495 0.037 0.082 100.0 100.0 0

d<2.7 38.11 0.089 0.55 3.89 8.48 0.016 0.022 23.5 10.7 20.0 25.1 27.3 49.7 1 NO o 2.7

d>4.3 19.58 0.222 7.09 15.45 17.59 0.026 0.009 30.3 70.8 40.9 26.8 22.3 10.2

1 1 000 1 00 0 1 1 00 0 1 1 00

100.0 0.144 1.96 7.40 12.87 0.022 0.016 100.0 100.0 0 oi o ON' o d<2.7 36.72 0.092 0.97 4.54 0.015 0.019 23.6 15.2 29.5 28.8 50.2 1 ON CN 2.7

-149+105 16.78 3.5

O O d S d NO d 3.9

d>4.3 21.76 0.234 13.55 28.37 13.58 0.021 0.011 35.4 74.5 56.5 24.2 25.0 17.9

1 1 00 0 1 00 0 1 1 00

100.0 0.144 3.96 10.93 12.21 0.019 0.014 100.0 100.0 100.0 0 0 00 1 o- rs m

A u 00 od od 10. 5 19. 9 u-i rs 23. 6 39. 4 24. 3 24. 1 46. 6 100. 0 100. 0 0 00 1 NO vO un rs oq <>

C u od m rs 10. 3 18. 5 19. 8 38. 0 26. 9 25. 9 27: 7 24. 1 100. 0 100. 0 0 00 1 0 00 1 0 00 1 vO vq m

M g r-^ 11. 0 21. 1 31. 2 33. 9 24. 2 31. 4 28. 7 26. 3 - 000 1

12. 3 d

11. 5 o-'

f> 72. 6 25. 1 66. 5 Distributio n % 100. 0 - 100. 0 0 00 1 a m ov rs Ov p O vO m un O^ u-i rn d

d 80. 3 25. 7 E 78. 4 100. 0 100. 0 Stot . -a oq p oq rn rs rs

z rs' 16. 6 vo rs 19. 2 30. 0 23. 5 49. 6 22. 5 42. 4 100. 0 100. 0 EC 100. 0 T 0.01 5 0.01 3 0.01 2 0.02 7 0.06 2 0.01 4 0.00 5 0.01 5 0.01 0 0.01 3 0.01 2 0.00 5 0.01 3 0.01 3 0.40 3 0.01 8

u A u (ppm ) 3 O u e- 0.02 5 0.02 0 0.02 5 0.01 6 0.01 6 0.01 3 0.01 5 0.02 1 0.04 3 0.02 1 0.20 7 0.02 2 0.01 6 0.01 9 0.01 3 0.01 7 S Cu % « J? en 11. 8 8.8 7 6.9 1 9.9 5 9.9 3 11.4 3 10.5 1 11.5 1 12.0 3 12.4 1 10.4 4

eu 10.8 7 10.9 5 s 12.4 7 12.6 6 12.5 8 Grad e WD un 7.3 1 7.1 6 9.5 7 6.5 3 18.8 9 16.0 9 14.4 7 12.5 7 12.6 3 19.5 3 15.6 1 15.7 3 17.8 6 37.1 2 33.1 3 2 Fe %

H O TT 1.7 5 1.4 4 1.1 4 7.4 8 3.6 0

Tf 3.7 6 5.3 6 4.8 5 3.4 5 2.9 8 5.1 4 5.8 6 2.2 8 16.2 6 14.6 8 u Stot. % 0.17 9 0.19 5 0.14 5 0.30 4 0.11 4 0.15 3 0.15 8 0.17 8 Ni % 0.14 2 0.19 3 0.18 1 0.13 4 0.10 5 0.09 0 0.11 0 0.17 1 3 cs Of , 1.1 9 1.7 6 1.3 9 1.7 2 6.4 6 9.4 3 100. 0 100. 0 36.9 4 32.4 9 21.1 7 31.3 2 H 28.7 5 27.4 0

m i d >4. 3 d >4. 3 d <2. 7 Clas s d <2. 7 Ov Densit y 3.5

Of .o 100. 0 16.9 5 21.6 3 27.8 3

r~ rn -74+3 7 -105+7 4 • Tot.(calc ) LO

CO

CM

CO

CO CO

T T U t/3 < 00

S 3 5» cn V cu Qi CO •*-> cn cn

CQ cu 3 cs H CO co

cn cu H JS cu

OQ

CO CO

S .S CO ro 0) CO .a 13 TJ a

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& g/ s 12. 1 7.8 6 2.4 2 0.5 5 0.2 9 0.1 6 0.7 9

+ I* 1.4 3 6.5 2 2.6 1 4.7 4 100. 0 19.5 0 oo to © 65.2 2 fN Q Ni % 0.23 1 0.27 1 0.28 5 0.28 5 0.36 9 0.25 9

2 ^ 9.4 8 11.7 1 H 10.8 4 11.5 8 12.5 2 14. 7 S % 7.6 1 2.8 5 4.3 5 4.2 5 4.4 6 4.1 9 9.3 1 16.5 8 18.2 6 19.5 7 15.4 9 Fe % 37.6 7 21.6 2 Fee d A u 0.04 3 0.02 5 0.02 8 0.03 8 0.04 0 0.02 0 0.03 1 (ppm ) g/ s 1.2 3 8.3 3 2.0 3 0.7 1 57.7 5 21.6 9 23.7 7 0 00 1 1.2 4 3.9 7 2.3 1 15.0 0 35.7 4 41.7 4

m V Tot . Clas s d>4. 3 d<2. 7

Densit y V 3.13 7 Siz e Class(um ) 138

90 -I -•- - - Test 1 80 •»-• Test 2 70 Test 3 60 Test 4

50 Test 5b

40 J, -•- • - Test 6

•+•-• Test7 30 .4.

20 -<«\

10

0 10 100 1000 Size (pm)

Figure C.6.5. Size Separation Partition Curves, SC-4

2.25 2.75 3.25 3.75 4.25 4.75 Density (g/cm3)

a -297+149 pm -149+105 pm -105+74 pm x -74+37 pm • +37 pm •+297 pm

Figure C.6.6. Partition Coefficient Curves for each Size Fraction Mass Yield versus Density) 2.25 2.75 3.25 3.75 4.25 4.75 Density (g/cm3) • 297 (jm • -297+149 pm -149+105 pm -105+74 pm X -74+37 pm • +37 pm

Figure C.6.8. Mg Recovery for each Size Fraction (Mg Recovery versus Density) 140

2.25 2.75 3.25 3.75 4.25 4.75 Density (g/cm3)

4> +297 pm • -297+149 pm -149+105 pm -105+74 pm x -74+37 pm • +37 pm

Figure C.6.9. S Recovery for each Size Fraction (S Recovery versus Density)

Figure C.6.10. Fe Recovery for each Size Fraction (Fe Recovery versus Density) 141

Appendix HE Flotation Test

Flotation Test on CVD Concentrate

Sample: CVD Concentrate (CVD Pilot Test 7) Sample weight: 500 g Sample Particle Size: Pgo = 160 [im

For flotation test the sample was ground to Pso - 100 |im.

Flotation Procedures and Conditions:

Conditioning time: 5 minutes pH: 10 pH Regulator: Soda Ash Collector: PAX (125 g/t) Frother: MIBC Rougher flotation time: 5 minutes

Table 1. Metallurgical Balance, SC-4

weight Grade Distribution % Material % Ni% Fe% Mg% S% Au(g/t) Ni Fe Mg S Au Concentrate 35.1 1.32 50.4 2.7 29.7 3.97 64.1 37.6 26.1 38.9 98.2 Tailing 64.9 0.40 45.2 4.1 25.2 0.04 35.9 62.4 73.9 61.1 1.8 Feed 100 0.72 47.0 3.6 26.8 1.42 100 100 100 100 100 (CVD Con)