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PENALTY ELEMENT SEPARATION FROM CONCETRATES UTILIZING

by Zachery Zanetell

A thesis submitted to the Faculty and the Board of Trustees of the Colorado School of Mines in partial fulfillment of the requirements for the degree of Master of Science (Metallurgical and Materials Engineering).

Golden, CO

Date:

Signed: Zachery A. Zanetell

Signed: Dr. Patrick R. Taylor Thesis Advisor

Golden, CO

Date:

Signed: Dr. Michael Kaufman Professor and Head Department of Metallurgical and Materials Engineering

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ABSTRACT

The copper that are currently being considered for development and processing are lower in grade and contain higher amounts of deleterious elements, which create difficulty in achieving a final copper concentrate that meets current restrictions. This presents increasing challenges to the process metallurgists during project development as well as to presently operating mines and mills. This thesis will focus on the separation of the deleterious elements, also known as penalty elements, mainly and from a copper concentrate using froth flotation techniques. The ability to separate penalty elements from copper concentrates will directly benefit companies by creating a final copper concentrate that will result in fewer financial penalties from smelter refineries.

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Table of Contents

Abstract ...... iii

List of Figures ...... vii

List of tables ...... viii

List of Equations ...... xi

Acknowledgements ...... xii

Chapter 1 Introduction ...... 1

Chapter 2 Literature Review ...... 3

2.1. Fundamentals of Froth Flotation ...... 3

2.2. Selective Oxidation of Arsenic during Flotation ...... 6

2.3. Depression of Arsenic Minerals in Flotation by Controlled Potentials ...... 7

2.4. Depression/Flotation of Arsenic Minerals ...... 8

2.5. Bismuth Flotation ...... 9

2.6. Characterization of Ores and Flotation Products ...... 11

Chapter 3 Experimental Procedures ...... 13

3.1. Testwork Samples ...... 13

3.2. Blending, Splitting and Storage of Testwork Samples ...... 13

3.3. Mineralogical Sample Preparation ...... 14 3.3.1. Preparation of Epoxy Mounts ...... 15 3.3.2. Surface Preparation of Epoxy Mounts ...... 16 3.3.3. Coating ...... 17 3.3.4. Liberation Analysis by Scanning Electron Microscopy ...... 17

3.4. Analytical Methods ...... 19 3.4.1. Four Acid Digestion ICP-260-X to determine Ag, Bi, Cd, Cu, Fe, Pb and Zn ...... 19

3.5. Analysis of Bismuth by Bi-ICP-MS ...... 20

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3.5.1. Standard Operating Procedure- Perkin Elmer NexION ICPMS 300D ...... 20 3.5.2. Perkin Elmer NexION ICP-MS 300D Method ...... 21

3.6. Chemicals ...... 21

3.7. Metallurgical Methods ...... 22 3.7.1. Flotation Procedure ...... 22 3.7.2. Particle Size Characterization – Screen Analysis ...... 23 3.7.3. Particle Size Analysis – Cyclosizer Analysis ...... 24

Chapter 4 Characterization of Copper and Testwork Feed Samples ...... 25

4.1. Analytical Characterization of Head Samples ...... 25

4.2. Mineralogical Characterization by XRD ...... 28

4.3. Mineralogical Characterization by MLA ...... 28 4.3.1. MLA Calculated Assay vs. Chemical Analysis ...... 29 4.3.2. Distribution of Penalty Element Minerals – Final and Rougher Concentrates ...... 30 4.3.3. Modal Mineralogy ...... 30 4.3.4. Mineral Liberation by Free Surface – Bismuth Minerals ...... 35 4.3.5. Mineral Association – Bismuth Minerals ...... 41 4.3.6. Penalty Element Transport – Flowsheet Distribution ...... 42

Chapter 5 Metallurgical Testwork Results ...... 44

5.1. Metallurgical Testwork Approach ...... 44

5.2. Baseline Metallurgical Testwork Results- Final Concentrate ...... 45

5.3. Depression of Copper Minerals – Final Concentrate ...... 50

5.4. Depression of Bismuth Minerals – Final Concentrate ...... 52

5.5. Baseline Metallurgical Testwork Results- Rougher Concentrate ...... 58

5.6. Depression of Bismuth Minerals – Rougher Concentrate ...... 61

5.7. Pre-aeration Screening Tests – Rougher Concentrates ...... 70

5.8. Summary of Successful Separation Tests – Final and Rougher Concentrates ...... 70

5.9. Confirmation Test – Rougher Concentrate ...... 70

Chapter 6 Conclusions and Recommendations ...... 74

References Cited ...... 77

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Appendix A ...... 80

Appendix B ...... 86

Appendix C ...... 90

Appendix D ...... 93

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LIST OF FIGURES

Figure 2.1: Classical flotation cell schematic (Zhengzhou ZY Machinery CO.,LTD n.d.) ...... 4 Figure 2.2: General flotation diagram ...... 5 Figure 3.1: Grinding and polishing sequence for Struers TegraPol-35 ...... 17 Figure 3.2: Carbon coating stub color with corresponding Ångstrom ...... 17 Figure 3.3: SEM/MLA Setup used in Testwork Analysis ...... 18 Figure 3.4: EDAX EDS Detectors used in Testwork Analysis ...... 18 Figure 4.1: Photomicrograph of liberated PbBiSbAgS in the final concentrate ...... 36

Figure 4.2: Photomicrograph of a partially liberated AgPbBiS3 particle in the final concentrate ...... 37 Figure 4.3: Photomicrograph of a partially liberated Bismuthinite particle in the final concentrate ...... 38

Figure 4.4: Photomicrograph of a partially liberated AgPbBiS3 particle in the rougher concentrate ...... 39

Figure 4.5: False color image of AgPbBiS3 particle ...... 40

Figure 4.6: Photomicrograph of a liberated AgPbBiS3 particle in the rougher concentrate ...... 40 Figure 4.7: Flowsheet diagram of penalty elements ...... 43 Figure 5.1: Cumulative copper grade recovery curves for baseline tests on final concentrate ...... 48 Figure 5.2: Cumulative bismuth grade recovery curves for baseline tests on final concentrate ...... 49 Figure 5.3: Cumulative copper grade recovery graphs for bismuth depression test series on final concentrate...... 55 Figure 5.4: Bismuth grade recovery graphs for bismuth depression test series on final concentrate ...... 56 Figure 5.5: Cumulative copper recovery vs. cumulative bismuth recovery for bismuth depression series on final concentrate ...... 57 Figure 5.6: Cumulative copper grade vs. cumulative bismuth grade for bismuth depression series on final concentrate...... 58 Figure 5.7: Cumulative copper grade recovery curve for rougher concentrate baseline tests ...... 59 Figure 5.8: Cumulative bismuth grade recovery curves for rougher concentrate baseline tests ...... 61 Figure 5.9: Cumulative copper grade recovery curves for rougher concentrate bismuth depression series ...... 65 Figure 5.10: Cumulative bismuth grade recovery curves for bismuth depression test series on rougher concentrate...... 66 Figure 5.11: Cumulative copper grade vs. cumulative bismuth grade for bismuth depression test series on rougher concentrate ...... 67 Figure 5.12: Cumulative copper recoveries vs. cumulative bismuth recoveries for bismuth depression test series on rougher concentrate ...... 68

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LIST OF TABLES

Table 1.1: Penalty Element Costs (Newmont Mining Corporation 2011) ...... 2 Table 3.1: Summary of Flotation Circuit Samples Received ...... 13 Table 3.2: Summary of SEM/MLA Samples Prepared ...... 15 Table 3.3: Summary of SEM/MLA Run Parameters ...... 19 Table 3.4: Chemical Reagents used in Metallurgical Testwork ...... 23 Table 3.5: Summary of Sample Screen Particle Size Analysis ...... 24 Table 3.6: Summary of Sample Cyclosizer Particle Size Analysis ...... 24 Table 4.1: Sample Head Analytical Characterizations ...... 26 Table 4.2: Assay-by-Size Head Analytical Characterizations ...... 27 Table 4.3: Results of XRD Analysis ...... 28 Table 4.4: MLA Calculated Assay vs. Chemical Analysis for Final Concentrate ...... 29 Table 4.5: MLA Calculated Assay vs. Chemical Analysis for Final Concentrate ...... 29 Table 4.6: Distribution of Penalty Elements in the Final and Rougher Concentrates ...... 30 Table 4.7: Modal Mineralogy of Five Copper Concentrator Circuit Samples ...... 31 Table 4.8: Modal Mineralogy and Elemental Distribution of Final Concentrate ...... 33 Table 4.9: Modal Mineralogy and Elemental Distribution of Rougher Concentrate ...... 34 Table 4.10: Liberation by Free Surface for PbBiSbAgS in the Final Concentrate ...... 35

Table 4.11: Liberation by Free Surface for AgPbBiS3 in the Final Concentrate ...... 36 Table 4.12: Liberation by Free Surface for Tetradymite in the Final Concentrate ...... 37 Table 4.13: Liberation by Free Surface for Bismuthinite in the Final Concentrate ...... 38 Table 4.14: Liberation by Free Surface for PbBiSbAgS in the Rougher Concentrate ...... 38

Table 4.15: Liberation by Free Surface for AgPbBiS3 in the Rougher Concentrate ...... 39 Table 4.16: Liberation by Free Surface for Tetradymite in the Rougher Concentrate ...... 41 Table 4.17: Liberation by Free Surface for Bismuthinite in the Rougher Concentrate ...... 41 Table 4.18: Mineral Association in Final and Rougher Concentrate Samples ...... 42 Table 5.1: Metallurgical Testwork Summary ...... 45 Table 5.2: Baseline Metallurgical Results for Final Concentrate ...... 46 Table 5.3: Summary of Metallurgical Copper Depression Screening Tests for Final Concentrate ...... 50 Table 5.4: Metallurgical Results for Final Concentrate- Copper Mineral Depression Series ...... 51 Table 5.5: Summary of Metallurgical Bismuth Depression Screening Tests for Final Concentrate ...... 52 Table 5.6: Metallurgical Results for Final Concentrate - Bismuth Mineral Depression Series ...... 53 Table 5.7: Baseline Metallurgical Results for Rougher Concentrate Flotation Tests ...... 60 Table 5.8: Summary of Metallurgical Bismuth Depression Reagent Tests for Rougher Concentrate...... 62 Table 5.9: Metallurgical Results for Rougher Concentrate Flotation Tests ...... 63 Table 5.10: Metallurgical Results for Rougher Concentrate Flotation Tests ...... 64 Table 5.11: Metallurgical Results for Rougher Concentrate Pre-aeration Flotation Tests ...... 69

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Table 5.12:. Metallurgical Results for Final and Rougher Concentrate Successful Pre-aeration Flotation Tests ...... 71 Table 5.13: Metallurgical Results for Confirmation of Successful Test Conditions on Rougher Concentrate ...... 73 Table A-1: Particle Size Analysis Final Concentrate ...... 81 Table A-2: Particle Size Analysis Rougher Concentrate ...... 82 Table A-3: Particle Size Analysis Rougher Scavenger Concentrate ...... 83 Table A-4: Particle Size Analysis Rougher Feed ...... 84 Table A-5: Particle Size Analysis Rougher Tails...... 85 Table B-1: Cyclosizer Data Final Concentrate ...... 87 Table B-2: Cyclosizer Data Final Concentrate 5 Minute Regrind ...... 88 Table B-3: Cyclosizer Data Final Concentrate 10.0 Minute Regrind ...... 89 Table C-1: Copper Mineral Liberation by Free Surface for Final Concentrate ...... 91 Table C-2: Copper Mineral Liberation by Free Surface for Rougher Concentrate ...... 92 Table D-1: Flotation Data Sheet Z10-6-1 ...... 94 Table D-2: Flotation Data Sheet Z10-6-2 ...... 95 Table D-3: Flotation Data Sheet Z10-6-3 ...... 96 Table D-4: Flotation Data Sheet Z10-8-1 ...... 97 Table D-5: Flotation Data Sheet Z10-8-2 ...... 98 Table D-6: Flotation Data Sheet Z10-8-3 ...... 99 Table D-7: Flotation Data Sheet Z10-8-4 ...... 100 Table D-8: Flotation Data Sheet Z10-8-5 ...... 101 Table D-9: Flotation Data Sheet Z10-8-6 ...... 102 Table D-10: Flotation Data Sheet Z10-8-7 ...... 103 Table D-11: Flotation Data Sheet Z10-8-8 ...... 104 Table D-12: Flotation Data Sheet Z10-8-9 ...... 105 Table D-13: Flotation Data Sheet Z10-8-10 ...... 106 Table D-14: Flotation Data Sheet Z10-8-11 ...... 107 Table D-15: Flotation Data Sheet Z10-8-12 ...... 108 Table D-16: Flotation Data Sheet Z10-8-13 ...... 109 Table D-17: Flotation Data Sheet Z10-8-14 ...... 110 Table D-18: Flotation Data Sheet Z10-8-15 ...... 111 Table D-19: Flotation Data Sheet Z10-8-16 ...... 112 Table D-20: Flotation Data Sheet Z10-10-1 ...... 113 Table D-21: Flotation Data Sheet Z10-10-2 ...... 114 Table D-22: Flotation Data Sheet Z10-10-3 ...... 115 Table D-23 Flotation Data Sheet Z10-10-4 ...... 116

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Table D-24: Flotation Data Sheet Z10-10-5 ...... 117 Table D-25: Flotation Data Sheet Z10-10-6 ...... 118 Table D-26: Flotation Data Sheet Z10-10-7 ...... 119

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LIST OF EQUATIONS

Equation 2.1: Ratio of concentration = F/C ...... 6 Equation 2.2: Percent Recovery = 100(c/f)(f – t)/(c – t) ...... 6 Equation 2.3: Metal Loss = 100 - Metal Recovery ...... 6 Equation 2.4: Weight Recovery = 100•C/F = 100•(f – t)/(c – t) ...... 6

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ACKNOWLEDGEMENTS

I would like to thank Dr. Ronel Kappes, John Gathje and Denise Doerr for their support and guidance throughout the entire thesis program; it was through their support and mentoring that made this thesis possible. I would also like to thank my advising committee members, Dr. Patrick Taylor and Dr. Corby Anderson for their help and accommodations throughout the Colorado School of Mines Master of Science program. The Newmont Metallurgical Services (NMS) team should also be recognized for their support and assistance throughout the course of this thesis. I also would like to thank Marc LeVier and Sevket Acar for their administrative support of my Master’s program.

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CHAPTER 1 INTRODUCTION

Copper is a well-known metal used in a variety of applications due to its high conductivity, recyclability and ability to resist corrosion. Within the last decade the mining industry has seen copper prices fall to record lows and most recently to record highs. Many of the current and future copper operations around the globe are processing lower grade copper ores at high tonnages. A potentially financially beneficial processing application to concentrate mined low grade copper deposits is through froth flotation. The ore is crushed and ground through a series of communition steps prior to feeding the material as slurry to a flotation circuit. Depending upon the mineralogy of the deposit the ore will be ground to fine particle sizes, P80 < 150 microns. Chemical additives adhere selectively to the slurry particles in flotation tanks where air is bubbled through the tanks creating a froth that can be collected. The chemical additives cause the copper bearing minerals to become hydrophobic and attach to the bubbles as they move up through the slurry. The copper minerals are then separated from the remainder of the slurry which is referred to as ; the copper minerals are collected in the froth. Depending on the flowsheet of the copper concentrator the copper minerals collected in flotation will be the final saleable copper concentrate, although several series of flotation circuits may be required to achieve the saleable final copper concentrate. In most cases the final copper concentrate that is to be sold to a smelter for also contains deleterious elements that degrade the copper concentrate. Deleterious elements are a growing concern for many copper concentrators requiring the sale of their copper concentrate. Due to the current and future demand for copper the environmental impact of mining is also a growing concern due to the elements that can become concentrated during copper concentration methods, specifically during froth flotation (Mular, Halbe and Barratt 2002). The increase in lower grade ores with deleterious elements is the motivation for separation techniques to meet these environmental regulations.

With the many existing copper concentrators around the globe and many more scheduled to become operational in the near future, separation of deleterious elements from the final saleable copper concentrate may be vital in order to avoid severe financial penalties from smelters. The deleterious elements in the copper mining industry are more commonly referred to as penalty elements. Penalty elements can be referred to as substituents that concentrate in the final copper concentrate reducing the overall grade and/or value of the copper concentrate for sale. Penalty elements are a broad term and can encompass a large variety of elements based upon a smelter contract between a specific copper concentrator and its buyer, the smelter. In the testwork presented the following penalty elements were considered; arsenic, bismuth, , , selenium and . Penalties suffered from the smelter to the copper concentrator are agreed upon prior to the sale in what is referred to as the smelter

1 contract. Table 1.1 presents the penalties that can be incurred from the smelter of the copper concentrates used in this testwork based upon the levels measured in the final copper concentrate.

Table 1.1. Penalty Element Costs (Newmont Mining Corporation 2011)

Penalty Limit/ Penalty Cost Element DMT1

Arsenic (As) 0.10% $ 5.00/0.1 % ( up to 0.5 % As) $ 11.00/ 0.1 % ( >0.5 % As) Bismuth (Bi) 200 ppm $ 4.00/ 100 ppm (up to 1200 ppm Bi) $ 6.00/ 100 ppm ( >1200 ppm Bi) Selenium (Se) 0.05% $ 5.00/0.01 % Se Antimony (Sb) 0.10% $ 4.00/0.1 % Sb Cadmium (Cd) 200 ppm $ 4.00/100 ppm Cd Lead (Pb) 1% $2.75/0.5 % Pb

1/ Dry Metric Tons

Copper concentrates can have significant penalty element concentrations depending on the nature of the ore source, specifically the mineralogy of the copper minerals. In this case the penalty elements are associated with minerals that contain elements of value causing the penalty elements to become concentrated in the final copper concentrate. As will be shown in later metallurgical results sections, the penalty tend to have the similar characteristics. Meaning that if a metallurgical separation can be applied to one penalty element the remainder of the penalty elements will tend to follow. Due to the varying nature of penalty element concentrations present in the final copper concentrate the main focus and study will be on bismuth. The main issue with bismuth during the refining processes is its ability to create a brittle final copper product degrading the value of the refined copper. The brittle copper product must be further treated in order to reach a purity of 99.99 % copper and the bismuth requires the smelter to follow hazardous waste restrictions in order to dispose of the element properly (Newmont Mining Corporation 2011).

As stated, copper concentrate penalties can vary significantly based upon the copper mineralogy of the ore sources and large financial penalties can be incurred from smelters. The combination of this significant penalty incurred with bismuth levels and the insufficient work that has been published about bismuth flotation separations in copper concentrates are the reasons for this study.

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CHAPTER 2 LITERATURE REVIEW

Literature surrounding the flotation and/or depression of penalty elements is scarce at best. A handful of papers were found that contained information on some sort of treatment of penalty element ores from around the world. The majority of the literature studied arsenic and the environmental effects caused by the presence of arsenic in certain ore bodies. The separation of arsenic in many different ore bodies from other value carrying has been well studied and documented. However, the literature which was found relevant usually summarized a process not related to the separation of arsenic from copper concentrates. In most cases arsenic was being separated from sulfide ore bodies, but the mineralogy of the arsenic was similar to that of the copper concentrates tested in this program deeming the information relevant.

As for the remainder of the penalty elements studied, the literature was difficult to find. A handful of papers were found on bismuth flotation and/or depression in sulfide ores, which like arsenic, did not show applications specific to copper concentrates. The mineralogy of bismuth in the majority of the papers found was similar to the mineralogy of the copper concentrate used in this test program deeming the papers relevant for inclusion in the literature review.

2.1. Fundamentals of Froth Flotation

Flotation today is the dominate mineral concentration method that is widely used in many applications of sulfide flotation. Flotation is primarily based on the interfacial chemistry of the mineral particles in . The differences in the interfacial chemistries of the minerals present in suspension are the cause for the separations that can be obtained through flotation. Most minerals in sulfide flotation are naturally hydrophilic, -loving, which require the use of chemical additives to alter the surface chemistry of the minerals making them hydrophobic (water-hating). The chemicals that are added are referred to as collectors, which selectively absorb to the minerals of interest providing the hydrophobic layer on the mineral surface required for separation. A variety of chemicals can be added to the slurry to enhance flotation separations that include; activators, depressants, and dispersants. Activators increase the selectivity by enhancing the collector absorption on the mineral surface. Depressants work opposite to activators by preventing the absorption of the collector on the mineral surface. Dispersants ensure that the mineral surface is free of the collection of fine particles and that collectors can adhere to the mineral surface. Frothers are also added to the flotation circuit to improve the dispersion of the bubbles within the pulp and to control/modify froth characteristics (Kelly and Spottiswood 1989).

Collectors are the first point at which separations can occur by the organic molecule or selectively absorbing to the mineral surface. For a successful absorption the collector must create a hydrophobic

3 layer on the mineral surface, which will cause the mineral to attach to the bubble and be collected in the froth. Collectors can come in many forms based on the chemistry that is required to alter the surface chemistry of the minerals targeted for collection. Two main collector groups exist today, anionic and cationic collectors. Anionic collectors contain bivalent groups, more commonly known to be xanthates and dithiophosphates. In general the longer the carbon chain in the collector the less selective the collector will be during sulfide flotation. Cationic collectors have a positively charged polar group with the hydrocarbon chain, commonly known as amines.

Flotation requires the partial liberation of the minerals in order for the reagents to be successful in absorbing to the surface. In general the flotation feed is ground to ~150 microns and fed to the flotation circuit at a density based on the requirements of the operation. Once the slurry is in the flotation cells it is thoroughly mixed to keep all particles suspended. Air is then finely dispensed throughout the pulp/slurry at a rate dependent upon the mineral system being utilized. Reagents can be added directly to the flotation cells, in conditioning tanks prior to flotation, or in the grinding circuit depending again on the mineral system being processed. Once the air has been added to the pulp the minerals that have had collectors adhere to their surfaces making them hydrophobic, attach to the bubbles and float to the top of the cell to be collected in the froth. Minerals can enter the froth in two ways, either by attachment of the hydrophobic particle to the bubble surface or through entrainment. Entrainment occurs when unwanted minerals are caught up in the water carried by the rise of the bubble layer through the pulp. Figure 2.1 illustrates a typical flotation cell and how minerals attach to bubble surfaces.

Figure 2.1. Classical flotation cell schematic (Zhengzhou ZY Machinery CO.,LTD n.d.)

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So by now it is apparent that flotation results in three products; feed (pulp), concentrate (collected froth), and tails (hydrophilic minerals). A variety of other products can be generated through flotation based on the mineral system and requirements of the operation to produce a final product. Figure 2.2 illustrates a general diagram of products and flotation designations that can result from a flotation circuit. The ore enters the flotation circuit as a pulp with a specified density. The first stage of flotation is referred to as rougher flotation that generates a rougher concentrate and rougher . The rougher tailings can be processed through scavenger flotation where additional reagents may be added or additional float time mat be added. The purpose of this stage is to scavenge the last remaining minerals of interest. Scavenger flotation may also be bypassed and the rougher flotation tails may be sent to a tailings dam or in many cases in gold flotation be processed through a leach circuit. The rougher concentrate and scavenger concentrate may be combined together and processed through cleaner flotation. Cleaner flotation may require a re-grind prior to flotation depending upon the mineral system being processed. In general cleaner flotation is operated at slurry densities that are significantly less than the density of the rougher and rougher scavenger stages. The cleaner tails can also be processed through a scavenger cleaner flotation stage similar to the scavenger stage discussed following rougher flotation stage. The concentrate collected during cleaner flotation is generally not combined with the cleaner scavenger concentrate. In general, the cleaner feed is reground prior to cleaner flotation to increase the liberation of the minerals of interest. The cleaner concentrate can be processed through many additional cleaner stages based upon the upgrading of the concentrate that is required.

Feed (pulp) Rougher Scavenger Tails Float Float

Rougher Scavenger Concentrate Concentrate

Cleaner Scavenger Tails Float Float

Final Scavenger Concentrate Concentrate

Figure 2.2. General flotation diagram

Flotation calculations can be generated based on the products observed in Figure 2.2 and some general information surrounding the circuit. Typically chemical analysis would be performed on each product and flowrates of each stream would be measured. Typically flotation calculations include the ratio of concentration, metal recoveries, metal losses, and weight recovery. The ratio of concentration is

5 calculated by taking the weight of the feed relative to the concentration and is shown in Equation 1 (Kelly and Spottiswood 1989).

Equation 2.1. Ratio of concentration = F/C (2.1)

The total weight of the feed (F) is divided by the total weight of the concentrate ©. Metal recovery calculations are probably the most important calculations in flotation and is expresses in Equation 2, where a percentage of the metal in the feed is recovered in the products from flotation (Kelly and Spottiswood 1989).

Equation 2.2. Percent Metal Recovery = 100(c/f)(f – t)/(c – t) (2.2)

Metal loss is different than metal recovery because this is the portion of the metal that is lost to the tailings of flotation and is calculated using Equation 3 (Kelly and Spottiswood 1989).

Equation 2.3. Metal Loss = 100 - Metal Recovery (2.3)

Weight recovery is the inverse of the ratio of concentration and is important to know. Equation 4 shows a typical weight recovery equation (Kelly and Spottiswood 1989).

Equation 2.4. Weight Recovery = 100•C/F = 100•(f – t)/(c – t) (2.4)

The calculations presented are generic and can be reworked based on the needs of the flotation circuit being assessed. All calculations presents are in no way inclusive, but show the basics in the calculation fundamentals of flotation (Kelly and Spottiswood 1989).

2.2. Selective Oxidation of Arsenic Minerals during Flotation

Arsenic in many applications can be an undesirable element that causes serious environmental problems during the treatment of concentrates during the process. Penalties imposed by smelters can be high causing significant financial implications to a project or operation. The separation of arsenic sulfides from non-arsenic sulfides has been well documented and studied throughout the mining industry. The difficulties in the during flotation are that the behaviors of the sulfide minerals are similar. Past studies have shown that the oxidation rates of the minerals studied at pH 11.0 follow the order of chalcocite> > enargite> covellite> chalcopyrite (Fornasiero 2001).

A separation technique to remove arsenic from non-arsenic sulfides requires the selective oxidation of the arsenic bearing minerals during flotation. A study conducted by Fornasiero in 2000 (Fornasiero

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2001) showed that arsenic sulfides could possibly be oxidized more than those of non-arsenic bearing sulfides causing the arsenic minerals to be depressed during flotation. A number of parameters were evaluated for oxidation, pH, and reagent levels on the different combinations of arsenic sulfides and non- arsenic sulfides. During baseline screening flotation tests, the separation of copper sulfide minerals of chalcocite, covellite or chalcopyrite from enargite or tennantite by flotation was found to be difficult. Better separation of the mineral systems described above was completed at pH 5.0 after the selective oxidation with peroxide (H2O2). Tests carried out at pH 11.0 after oxidation with H2O2 required the addition of ethylene diamine tetra-acetic acid (EDTA) to remove the oxidation of non-arsenic minerals prior to the addition of the collector. The study also concludes that an extensive conditioning time was required prior to flotation to oxidize the arsenic sulfides, thirty minutes.

The rate of oxidation of arsenic minerals is of critical importance in the understanding of their flotation behavior which can lead to the successful separation from non-arsenic bearing sulfides (Ma 2009). According to Ma 2009, enargite can display natural floatability under oxidizing conditions and non- floatability under reducing conditions with collector-less flotation parameters. However, many arsenic minerals can be readily floated using different chemical compositions of xanthates (salt with the formula - + + + ROCS2 M (R = alkyl; M+ = Na , K ) (Cytec Mineral Processing Chemicals 2013). Since many operations require the use of xanthates to recover sulfide values the floatability of arsenic sulfides with the use of xanthates poses a problem in the separation process. As noted by Ma in 2009 the flotation recovery behavior of sulfide minerals is attributed to the presence of the layer preventing the of the collector to surfaces of the sulfide minerals. It is also suggested that in order to separate enargite from chalcocite the sample must be conditioned with H2O2 (Ma 2009). In this case, the chalcocite was oxidized to a greater extent than the enargite allowing the chalcocite to be depressed and enargite to be recovered. It was also found that arsenic sulfide oxidation rates can differ vastly between minerals causing the process to become subject to the mineralogy of the ore source.

Separation of arsenic sulfide minerals from other sulfide minerals can be achieved using methods of pre-oxidation of the flotation pulp, depressants/collectors, and the control of the slurry potential. All processes have their applications based upon the mineralogy of the slurry and the requirements of smelter contracts (Fornasiero 2001) (Ma 2009).

2.3. Depression of Arsenic Minerals in Flotation by Controlled Potentials

A study conducted by Subramanian in 2007 evaluated the non-selective depression of all sulfides followed by the re-activation of the pyrite with a suitable collector can separate arsenic minerals from other sulfides at controlled potentials (Subramanian 2005). The flotation process produces a concentrate of low mass and high recoveries of sulfide minerals containing gold. The concentrate is then processed through which recovers about 50 % of the gold. Mineralogy of the concentrate revealed

7 that significant portions of the gold are associated with gersdorffite (NiAsS) and arsenopyrite (FeAsS). The depression of the sulfides can be carried out by conditioning the concentrate with potassium permanganate at a controlled potential for a timed period. The redox potential of the concentrate was held at 450 mV using platinum versus saturated calomel electrode by the addition of the potassium permanganate. A total of 10.0 minutes was used for conditioning time with the potassium permanganate followed by the addition of A404 (Mercaptobenzothiozole) to float the pyrite in the first flotation stages. Optimum results were shown to first be the collection of the talcose minerals leaving a sulfide rich product behind. Followed by selective activation of the pyrite by AP404 leaving the arsenopyrite minerals behind. The tailings from the tests were shown to be arsenopyrite rich and the pyrite concentrate exhibited the best cyanidation results after re-grinding.

Another study conducted by Bruckard that was published in 2007 took a synthetic mixture of metallic arsenic and quartz to test the floatability of metallic arsenic over different ranges of pH and Eh using ethylxanthate (KEX). Since metallic arsenic has no real floatability many different parameters had to be examined. Flotation tests were conducted over a wide range of conditions, pH 8-12 and Eh range of -500 mV to + 500 mV (SHE). It was found that metallic arsenic will float over a pH range of 5-10 when KEX is added to the flotation pulp, but significantly lower when the pH is raised above 11 (Bruckard 2007).

Control of slurry potential (Eh) was also studied in many cases and has been shown to offer benefit to the separation process. For example, it was found that the flotation of enargite is Eh dependent in the presence of xanthate at pH 10.5. The recovery of enargite was shown to be 100% at Eh levels between 0.15 and 0.27 V (standard hydrogen electrode), but decreased at levels outside of the range (Jamie and

Cifuented, 1995; Kantar, 2002). Enargite-chalcopyrite separation through Eh control has been studied but disagreement in the flotation parameters exists (Ma 2009). Two optimal conditions were found from the study to float metallic arsenic by Bruckard in 2007; at pH 6 metallic arsenic will float well within the Eh range of +125 to +275 mV (SHE), but seems to drop off at Eh’s above +375 mV and below +125 mV (SHE). Metallic arsenic was also shown to float at pH 10 under all potentials until +225 mV (SHE), but completely dropped off at +300 mv (SHE). Since metallic arsenic was successfully separated over a wide range of conditions using a simple collector the possibility of separating arsenic from other base metals should be possible (Bruckard 2007) using controlled potentials and the use of Xanthate collectors such as KEX.

2.4. Depression/Flotation of Arsenic Minerals

A considerable amount of research has been done on the separation of arsenopyrite from pyrite under different flotation parameters (Dai 2005). Arsenopyrite has been shown to be depressed with the addition of such chemicals as an ammonium salt- oxide or permanganate. However, there is little literature on the depression of arsenic minerals from the flotation of pentlandite (Fe,Ni)9S8. The reason for the need to separate arsenic from the Ni-Cu concentrate is that arsenic modifies the morphology of

8 the matte which to a decrease in the grain size of the copper sulfides. The high arsenic can also make the concentrate unsuitable for further chemical and metallurgical processes.

Since the majority of arsenic depression has been completed with regards to arsenopyrite a literature search was conducted to explore work which considered the rejection of other arsenic minerals. David and Quast 1991 tested several reagents for the depression of loellingite (FeAs2). Several reagents, that included, potassium cyanide, ferrous sulfate, sodium hypochlorite and Cytec flotation depressants 610 (polymeric depressant.

Literature was found on the depression of arsenic in pentlandite flotation using oxidative conditioning. Several experiments surrounding flotation parameters were studied in order to selectively oxidize the arsenic minerals during Cu-Ni flotation. In the first method, rougher and scavenger stages were followed by the cleaning of the combined rougher and scavenger concentrates. A mixture of magnesium chloride, ammonium chloride and ammonium hydroxide (MAA) was added to the rougher stage prior to the addition of the collector after the slurry had been pre-aerated for 30.0 minutes. It was shown through several tests that the MAA alone could not depress the arsenic minerals during flotation without the pre-aeration step. The tests using this procedure also showed that pre-aeration decreased the floatability of pyrrhotite, but the addition of MAA increased pyrrhotite flotation decreasing the grade of the concentrate.

A condition was sought to eliminate the effect of the pyrrhotite by removing the mineral prior to flotation using . After magnetic separation, triethylenetetramine and sodium sulfite (TSS) was also added to depress any remaining pyrrhotite in the flotation of the Cu-Ni concentrate (Dai 2005). Through the investigation of this flowsheet option, the combination of magnetic separation with the addition of TSS was shown to decrease the amount of pyrrhotite in the concentrate. TSS was also shown to be successful in the depression of arsenic containing minerals in Cu-Ni flotation (Dai 2005). The last set of flowsheet options investigated the cleaning of the scavenger concentrates to remove additional arsenic. The magnetic and scavenger concentrates were combined and reground prior to cleaner flotation. Addition of TSS or MAA in the cleaner stage greatly reduced arsenic recovery to the cleaner concentrate, the addition of TSS or MAA after pre-aeration offered additional rejection of arsenic minerals.

2.5. Bismuth Flotation

Certain complex ores that are found around the world, often contain penalty elements in association with the gold and values. Bismuth can be one of these penalty elements. In recent years bismuth has begun to move from a penalty element to a valuable metal in some cases, depending on mineralogy and content. Conventional methods of gold-bismuth processing have included gravity, gravity/flotation or

9 other hydrometallurgical combination methods. Bismuth-gold concentrates produced have been, in the past, cyanide leached to recover the gold and silver leaving the bismuth in the residue. The bismuth can also be recovered from the residues by leaching with (Khmelnitskaya 2007).

Sulfide minerals present in the ore characterized for a study conducted by Khmelnitskaya in 2007 were pyrite (FeS2), (PbS), bismuthinite (Bi2S3), bornite (Cu5FeS4), chalcocite (Cu2S), pyrrhotite

(Fe1-xS) and arsenopyrite. Liberated gold occurs as native gold or in association with bismuthinite, bismuth lead sulfosalts, pyrrhotite, arsenopyrite, magnetite, pyrite or quartz. Two pilot plant options were tested in order to increase the recoveries of gold, silver, and bismuth. As was shown in the testwork conducted, gravity/flotation followed by the hydrochloric leach to recover the bismuth and then cyanide leached to recover the gold and silver was found to be the most beneficial. Shown from the study conducted by Khmelnitskaya, bismuth can be separated from sulfides by leaching with hydrochloric acid solution. Copper is generally leached in acid systems which would not allow a separation of copper and bismuth if both existed in an acidic leach solution.

Another study conducted by Newcombie that was presented at The Australian Institute of Mining and Conference in 1984 evaluated the early record of processing copper ore at Warrego that produced a concentrate that contained 1.3% bismuth, smelting was accomplished in a flash furnace to eliminate the bismuth (Newcombie 1984). Geological evidence began to suggest that bismuth levels in the ore would continue to increase leaving the smelting process unable to eliminate the bismuth needed to make a saleable concentrate. A variety of flotation process options were investigated including the use of sodium cyanide and cyanide as bismuth depression reagents, but neither yielded a process that was successful in the rejection of the bismuth from the final concentrate to the Warrego concentrator. Perhaps the mineralogy of the copper and bismuth minerals was the reason this ore could not be separated under the flotation conditions investigated, bismuth liberation also seemed to be an issue.

A handbook issued by Cytec Industries offers some guidance on some flotation conditions that might aid in the separation of copper from some penalty elements (Thomas 2010). Since the majority of the bismuth mineral studied in this test program; were associated with a lead-silver-sulfide mineral as will be discussed later, many reagent options for lead silver depression/flotation reagents could be considered. The goal of the testwork was also to produce a copper concentrate, which meant that an elevated pH system had to be considered in order to preserve copper grade. This eliminated a decent amount of lead sulfide reagents with the exception of two polymeric reagents, Cytec 7262 and 7261A. Both reagents are labeled as depressants with the primary acting chemical source being polyacrylamide (Thomas 2010). Based on the information provided in the Cytec Chemical Handbook and conversations with Cytec experts (Maes 2011) both chemicals were selected to be used in the testwork that will be presented in later sections. The primary use for both chemicals was said to be for lead depression, which would aid in

10 the depression of the lead in the samples received and depress a portion of the bismuth. However, most recently they have both been shown to aid in the depression of penalty elements (Cytec Mineral Processing Chemicals Technology 2013).

2.6. Characterization of Ores and Flotation Products

To begin to identify a solution to a mineral processing problem mineralogical characterization needs to be completed on samples identified for testwork. Mineralogy can hold the key to unlocking the solution to a problem based upon the characteristics of the ore can include mineral locking, liberation and association. It was important to understand the mineralogy in the samples that were tested to identify how bismuth and copper were associated which might lead to a successful separation. In a study by Kouzmanov conducted in 2005 bismuth and tellurium bearing minerals were shown to be in association of copper and gold mineralized deposits. The Elshitsa and Radka deposits were examined for the mineralogy of these bismuth and tellurium bearing minerals. The deposits are hosted by dacitic volcanic and sub volcanic rocks with two types of ores, massive pyrite and Cu-Au sulfide ore. Both deposits show bismuth (Bi) and tellurium (Te) bearing minerals to form in small inclusions with the main ore-forming minerals of economic value, gold and copper minerals. The most common Bi and Te bearing mineral is tetradymite forming tabular idiomorphic crystals up to 100 microns in length. The samples identified for testwork displayed various amounts of bismuth that was locked within a bismuth telluride, tetradymite

(Bi2Te2S). Although the study by Kouzmanov didn’t offer separations of the telluride-bismuth mineral from copper sulfides, it did however explain the geological occurrences of tetradymite. The study also accurately displayed how mineralogical tools can aid in identifying the occurrence of bismuth within samples, which could potentially lead to separations.

It is important in any mineral processing application to understand the mineralogy of the ore in order to identify a successful separation process. NMS summarized data obtained from a scanning electron microscope (SEM) and analyzed with a Mineral Liberation Analyzer (MLA) on copper concentrates from operations (Kappes 2007). In order to separate values from penalty elements, the mineralogy of the deposit must first be characterized in order to understand their association, liberation characteristics and whether the sample received is appropriate for further testwork. A test program at NMS was initiated in order to reduce the arsenic in the final copper concentrate produced. Prior to testwork being initiated, head ore characterization needed to be completed in order to understand the mineralogy of the deposit in order to develop a successful separation technique. Through SEM/MLA sulfide mineral occurrences were characterized in the sample to include arsenic and copper mineralogy.

Several powerful tools can be used for mineralogical characterization of a deposit to include elemental distribution, grain size analysis, modal mineralogy, mineral liberation data, and mineral association. Elemental distribution shows the elemental breakdown of the elements present in the

11 sample corresponding to the minerals that are present. Grain size analysis will indicate the liberation size the sample must be ground to in order to have a successful separation of individual elements, which is used in conjunction with mineral liberation data. Modal mineralogy is a tool that allows the metallurgist to identify the minerals that are present in the sample at the various levels of occurrence. Mineral associations data is important to the testwork that will be presented because of the potential of arsenic to be associated with copper minerals which can help the metallurgist understand develop an experimental separation procedure. For example, if a large amount of the arsenic is associated with copper a liberation issue could be the source of a problem and regrinding of the sample followed by flotation tests might be performed. R. Kappes, D. Brosnahan and J. Gathje in 2007 successfully demonstrated that SEM/MLA characterization of laboratory flotation products can provide valuable information to develop testwork parameters. Understanding the mineralogy of a deposit can provide a metallurgist with powerful knowledge to tackle many of the separation processes they are faced with.

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CHAPTER 3 EXPERIMENTAL PROCEDURES

This chapter will discuss the experimental procedures used for the preparation of the samples received for testing, mineralogical and analytical characterization methods used to characterize penalty elements, and copper concentrate flotation procedures.

3.1. Testwork Samples

The copper concentrates that were used for this research were obtained from an operating copper concentrator in North America. Five flotation circuit samples were received; a final copper concentrate, a rougher concentrate, a rougher scavenger concentrate, a rougher feed, and a final tails sample. Table 3.1 summarizes the samples received and the corresponding weights. All five samples were received as wet filter cakes to limit the oxidation prior to penalty element separation testwork.

Table 3.1: Summary of Flotation Circuit Samples Received

Rougher Final Copper Rougher Rougher Sample ID Scavenger Rougher feed Concentrate Concentrate Tails Concentrate NMS Lot No. 110322-1 110322-2 110322-3 110322-4 110322-5

Wet Weight (kg) 21.5 7.6 5.4 5.4 3.5 % Moisture 23.3 11.0 15.6 17.8 20.1 Total Dry wt. (kg) 16.5 6.7 4.5 4.4 2.8

A variety of samples were requested and received as outlined in Table 3.1 in order to characterize the penalty elements in the flotation circuit at the copper concentrator. The samples were needed in order to be analyzed for penalty elements to create a balance showing where the penalty elements concentrated in the flotation circuit. The samples designated as rougher and final copper concentrates (110322-2 and 110322-1) were prepared and used for froth flotation testing in this study. The rougher and final concentrates were selected for penalty element separation testwork due to the potential ease of implementation of a separation process if one were to be developed. The remaining products were used in conjunction with the concentrates to provide a balance illustrating the deportment of the penalty elements through the flotation circuit. The subsequent sections will describe in detail the procedures and methods used for each aspect of sample preparation, mineralogical characterization, and analytical characterization of the samples (Kennedy 1990).

3.2. Blending, Splitting and Storage of Testwork Samples

The five samples received were shipped in separate five gallon plastic buckets labeled by their product designations. All samples were wet and contained filter papers with appropriate labels for easy

13 sample identification. The samples were each weighed in order to determine the original weight as received. A sub-sample from each of the five sample buckets was taken (grab sample) and dried in order to determine the moisture content of each of the samples. Table 3.1 summarizes the moisture content of each of the samples received. Once the moisture content of each the sample was determined the sample weights could then be calculated. The five samples were re-slurried with tap water, wet rotary slurry split, and each split was filtered individually. A split from each of the five samples was used for particle size characterization, mineralogical characterization, and analytical characterization (Smith 1973).

Since penalty element separation testwork would focus on the final and rougher copper concentrates additional splitting was completed. The two individual concentrates were re-slurried in order to retain the homogeneity of the samples. Once re-slurried the samples were then wet rotary slurry sampled in order to generate ~240 g flotation charges for each concentrate product to use for flotation testwork. Each of the prepared charges were filtered and sealed in vacuum packages for preservation. The remaining three samples were vacuum filtered as individual bulk samples should the need for future testwork arise. Once each of the samples was vacuum filtered, and vacuum packaged they were stored in a refrigerated cold storage location for future testwork. The reason the samples were refrigerated and not frozen was to preserve the samples and limit the oxidation of the minerals. Samples that are frozen may also have the crystal grain structure destroyed.

3.3. Mineralogical Sample Preparation

Particle size analysis was performed on each of the five samples to determine the P80, the size at which 80 % of the particles by mass are smaller than that particle size. The samples that were retained on each of the sieves, explained in the Particle Size Characterization Section 3.7.2, were used for Scanning Electron Microscopy (SEM) followed by Mineral Liberation Analysis (MLA).

The screen fractions from the particle size characterization were kept separate for each sample for mineralogical characterization. It was determined that three size fractions for each of the samples would be used for SEM/MLA based on the weight of sample retained in each size fraction. The size fractions for each of the five samples included; 65 x 270 mesh (208 x 53 microns), 270 x 500 mesh (43 x 25 microns), and -500 mesh (> 25 microns). Each of the screen fractions for each sample were combined and blended within their respected sample. Table 3.2 summarizes the breakdown of the screen fractions used for SEM/MLA analysis.

The determination of the size classifications was based upon the amount of weight of each size class retained for each sample. Sufficient sample had to be in each size class for mineralogical and analytical work to be completed.

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Table 3.2: Summary of SEM/MLA Samples Prepared

Rougher Final Copper Rougher Rougher Sample ID Scavenger Rougher feed Concentrate Concentrate Tails Concentrate NMS Lot No. 110322-1 110322-2 110322-3 110322-4 110322-5 Mesh Microns µm grams retained 65 208 0.0 1.0 0.0 5.4 24.7 65 x 100 147 0.0 2.1 2.0 8.6 29.2 Fraction 1 100 x150 104 3.6 4.6 3.5 7.6 21.9 65 x 270 150 x200 74 9.4 10.3 8.5 8.2 21.2 mesh 200 x270 53 11.9 13.8 9.4 7.1 17.4 Fraction Total 24.9 31.8 23.4 36.9 114.4 270 x325 43 5.2 5.9 4.0 2.6 7.1 Fraction 2 325 x 400 38 4.4 5.6 2.4 1.0 4.8 270 x 500 400 x 500 25 12.6 14.1 8.1 6.8 13.3 mesh Fraction Total 22.2 25.6 14.5 10.4 25.2 Fraction 3 -500 < 25 61.7 74.1 48.4 35.6 85.7 -500 Fraction Total 61.7 74.1 48.4 35.6 85.7 mesh

Total 155.9 188.9 124.2 130.2 364.9

P80 59 62 67 135 169

Subsequent to blending the sample retained for each size fraction, was split into half. One half of each fraction was used for SEM/MLA analysis and the other half was pulverized and submitted for chemical analysis. This procedure was followed for each of the 15 sample fractions. The assay split from each sample was pulverized with a ring and puck until the sample was fine, with a powder like (flour) consistency. The assay splits were submitted in a Laboratory Information Management System (LIMS) for chemical analysis at NMS. The SEM/MLA splits were also submitted in LIMS, but remained un- pulverized to retain mineral texture.

3.3.1. Preparation of Epoxy Mounts

A total of 15 samples were submitted for mineralogical characterization through SEM/MLA and epoxy mounts were made for each of the three size fractions for each of the five samples received. Each size fraction was weighed into an individual cup for epoxy mounting. Samples that were greater than 325 mesh had 4.0 grams of sample weighed into the individual epoxy mounting cups and were mounted perpendicular. The reason for the perpendicular mount was due to the settling nature of the particles in the coarse screen fractions, those > 325 mesh. The original epoxy mount had to be cut and mounted perpendicular to expose the settling layer of the coarse fractions for the new epoxy mount. Samples that

15 were less than 325 mesh had 1.0 gram of sample and 0.3 grams of carbon powder weighed into their individual cups (Doerr, et al. 2009). Next the samples were moved from the weigh bench to a ventilated hood to pour the epoxy liquid into each of the 15 epoxy mounting cups. In order to have enough time before the epoxy hardened, five cups were poured at a time. 11.0 grams of epoxy was mixed for each of the samples, which translated to 25 parts resin and 3 parts of hardener. The resin was added to the hardener and mixed for approximately 1.0 minute. Next the resin/hardener mixture was transferred to a stir plate for 5.0 minutes to ensure homogeneity of the epoxy. Upon conclusion of the five minutes the epoxy was removed from the stir plate and allowed to set for 5.0 minutes. Once the epoxy was completely mixed, approximately 11.0 grams of epoxy was dispensed into each of the sample cups. Using a wooden stick, the epoxy was then mixed with each of the samples in the cups. Once mixed thoroughly, the sample/epoxy mixture was transferred to an individual mold for each sample. The poured molds were then transferred into an for 10.0 minutes and then allowed to harden at room temperature for 1 hour. Once the mounts were semi-hardened LIMS labels were then added to each of the samples for identification purposes. Each sample was then left to fully harden before additional epoxy was added to each mold to cover the sample identification labels. After allowing the sample epoxy molds to harden with the additional epoxy, the samples were then released from the molds and transferred to the oven for 10.0 minutes. Once the samples were completely hardened they were moved to the polishing station.

3.3.2. Surface Preparation of Epoxy Mounts

Each of the epoxy mount samples was processed through a grinding stage prior to polishing to obtain a surface that was damage free. Two stages of grinding, Plane Grinding (PG) and Fine Grinding (FG) were used to prepare the epoxy mounts (F and C 1984). Plane grinding is referred to as the coarse grinding of the epoxy mount in which the surface of the epoxy is ground off to expose the surface of the minerals in the mount. Next the epoxy mounts are processed through fine grinding in order to further refine the surfaces of the minerals that were exposed during plane grinding.

The fifteen epoxy blocks were polished using Struers polishing instruments, TegraPol-35, to further expose a fresh mineral surface. Polishing was carried out using polishing cloths with a polycrystalline diamond suspension and lubricant/extender (water soluble). Fine polishing polycrystalline diamond consists of numerous minute crystallites that are used for high material removal while producing a small scratch depth. Lubricant/extender helps to enhance diamond performance and material removal on the epoxy mount.

Each of the epoxy mounts was polished with the following grit MD-Piano paper with the sequences shown in Figure 3.1. After polishing the samples were inspected individually under a microscope for

16 flaws in the finished epoxy mounts. Once each epoxy mount had successfully been screened for flaws the samples were ready for the next step, carbon coating.

Figure 3.1: Grinding and polishing sequence for Struers TegraPol-35

3.3.3. Carbon Coating

Each of the 15 epoxy mounts were carbon coated to ensure electrical conductivity at the surface to prevent the accumulation of electrostatic charge as the final stage for the polished epoxy mounts prior to SEM/MLA analysis. The samples were carbon coated using an EMS (Electron Microscopy Sciences) instrument. A brass stub was used as an indicator of the carbon coat thickness to insure optimum thickness, which is between 200 and 250 Å. Figure 3.2 illustrates the brass stub color and corresponding Ångstrom. The fifteen samples were carbon coated to 250 Å for the SEM/MLA runs at NMS.

Figure 3.2: Carbon coating brass stub color with corresponding Ångstrom

3.3.4. Mineral Liberation Analysis by Scanning Electron Microscopy

The SEM used for testwork was a FEI Quanta 600 SEM with 2 EDAX 10 mm2 Apollo XP silicon drift detectors (SDD) for energy-dispersive X-Ray spectrometry with ZAF corrections (Z=density, A=atomic number, and F=fluorescence). The SEM uses Mineral Liberation Software (MLA) version 3.1 accompanied with EDAX genesis software for EDS measurements for the bean control of the MLA measurements. Figure 3.3 shows the setup of SEM/MLA utilized in this test program and Figure 3.4 illustrates the EDAX/EDS detectors for the MLA.

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Figure 3.3: SEM/MLA Setup used in Testwork Analysis

Figure 3.4: EDAX EDS Detectors used in Testwork Analysis

The five samples presented in this thesis were analyzed using an SEM with mineral liberation software to provide information on the occurrences of the penalty elements, in the received samples. The

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SEM is accompanied by Mineral Liberation Analysis (MLA) software V3.1 and EDAX Genesis software for EDS measurements as well as beam control for MLA measurements.

The MLA method used to analyze the samples was extended back scatter electron liberation analysis (XBSE). XBSE is the more advance liberation method in which back scatter electron (BSE) images are collected and segmented to delineate the mineral grain boundaries in each particle that is scanned by the SEM. The minerals grains are then analyzed using an x-ray to generate the chemical composition of the particle. Particle maps are generated from particle segmentation data and x-ray spectra to be analyzed.

After each of the samples was mounted, polished and carbon coated each mounted sample was run on the MLA to generate mineralogical data. A range of parameters needed to be set in order to accurately analyze each sample. Fortunately previous samples from the copper concentrator have been analyzed with this SEM/MLA system and the majority of sample parameters are known. The parameters include BSE gray level, brightness, contrast, magnification, spot size, frames, particles processed and elapsed time. Table 3.3 summarizes the parameters used in analysis of the samples.

Table 3.3: Summary of SEM/MLA Run Parameters Gray Particles Elapsed Sample ID Sample Name Brightness Contrast Magnification Spotsize Frames Level Processed Time (mins)

Z10-4-1 +270 RO FEED 240 46 42 275 6.6 154 15018 138.6 Z10-4-1 270 X325 RO FEED Rougher Feed 244 46 42 350 6.6 392 15893 146.2 Z10-4-1 -500 RO FEED 241 46 42 700 6.6 322 16010 126.3 Z10-4-1 +270 FC 244 46 42 275 6.6 63 10146 62.3 Z10-4-1 270 X325 FC Final Concentate 247 46 42 350 6.6 255 10814 87.3 Z10-4-1 -500 FC 242 46 42 700 6.6 208 10680 73.4 Z10-4-1 +270 RC 240 45 42 275 6.6 76 10088 61.3 Z10-4-1 270 X325 RC Rougher Concentate 249 46 42 350 6.6 364 10177 113.3 Z10-4-1 -500 RC 242 46 42 700 6.6 200 10402 81.4 Z10-4-1 +270 RC SVC 242 45 42 275 6.6 81 10083 68.5 Z10-4-1 270 X325 RC SVC Rougher Scav Concentate 247 46 42 350 6.6 425 10142 128.3 Z10-4-1 -500 RC SVC 240 46 42 700 6.6 260 10567 87.1 Z10-4-1 +270 RT 240 46 42 275 6.6 293 20560 263.6 Z10-4-1 270 X325 RT Rougher Tail 244 46 42 350 6.6 518 20700 181.5 Z10-4-1 -500 RT 248 46 43 700 6.7 434 21115 174.5

3.4. Analytical Methods

Methods for the analytical results as well as the operation of the instruments used to determine the analytical results from feed and flotation analysis are explained in the subsequent sections.

3.4.1. Four Acid Digestion ICP-260-X to determine Ag, Bi, Cd, Cu, Fe, Pb and Zn

A four acid digestion method was used to determine the amount of total metals in heads and flotation products for this test program. First, each solid sample was ring and puck pulverized to insure homogeneity until fine flour like powder was obtained. The samples were then submitted to the analytical lab using LIMS for chemical analysis. The samples are weighed into a 250 mL Teflon beaker where acids are added in a specific order to digest the solid samples. The digestion takes place on a hot plate

19 in a fume hood until the sample is dry. The residues are then re-dissolved in hydrochloric acid and then transferred to 50 mL Digitubes.

The specific analytical procedure used for analysis is outlined below:

 Weigh 0.5 g of sample in to a labeled 250 mL Teflon beaker.  Acids are dispensed in the following order:  10 mL water

 10 mL NNO3 (Nitric Acid 16M)

 10 mL HClO4 (Perchloric Acid 11M)  5 mL HF (Hydroflouric 12M)  Swirl Teflon beaker with acids and place on hot plate with a Teflon watch , leave on hot plate for 30 minutes. After 30 minutes rinse each watch glass into the corresponding beaker and remove the watch glass.  Remove Beakers and cool  Add 10 mL of de-ionized water and then 10 mL of HCl ( Hydrochloric acid 12M)  Return beakers to hot plate, heat uncovered for 5.0 minutes  Transfer the sample from the Teflon beaker to a 50 mL Digitube. Allow sample to cool and then add 0.500 mL d-Tartaric Acid (10 %).  Bring to 50 mL volume with DI-water in Digitube  Allow sample to settle overnight  Run sample on ICP

3.5. Analysis of Bismuth by Bi-ICP-MS

This method was used to determine the amount of bismuth (Bi) in the head and flotation products that were submitted to the analytical laboratory. Due to low concentrations of Bi in the samples tested (particularly tails products); Bi-ICP-Mass Spectrometry (MS) was the method of choice. The limits of detection for the samples tested in this program ranged between 50 ppb and 1 ppm. The samples were digested using the four acid digestion process described in Section 3.4.1 and analyzed using the ICP-MS procedure described in Section 3.5.1.

3.5.1. Standard Operating Procedure- Perkin Elmer NexION ICPMS 300D

The Perkin Elmer NexION ICP-MS 300D is used for the sub-ppm analysis of the aqueous samples from the digestion process explained in Section 3.4.1. The Perkin Elmer NexION ICP-MS 300D is a combination of inductively coupled plasma with quadrupole mass spectrometer. The ICP-MS uses the ability of the argon ICP to efficiently generate singly charged from the elemental species within a sample. These ions are then directed into a quadrupole mass spectrometer. The function of the mass

20 spectrometer is similar to that of the monochromator in an AA or ICP emission system. However, rather than separating light according to its wavelength, the mass spectrometer separates the ions introduced from the ICP according to their mass-to-charge ratio. Ions of the selected mass/charge are directed to a detector which quantitates the numbers of ions present.

The NexION ICP-MS 300D model is equipped Universal Cell Technology; DRC (Dynamic Reaction Cell) and KED (Kinetic Energy Discrimination). The DRC uses reactions to reduce interfering ions. Gas inlets pressurize the reaction chamber with a low flow of reaction gas, such as , methane, , or other gasses. The reaction gas is selected based on its predictable ability to undergo a gas phase chemical reaction with the interfering species and remove the interference. Interference removal can occur through various processes, including collisional dissociation, electron transfer, proton transfer, and oxidation. A common interference is the overlap of Argon- Oxygen with at mass 56. By using a reactive gas, such as ammonia, the gas will react with the argon-oxygen compound to form ground state atoms of argon and oxygen. Gasses that are at ground state are not stable within the quadrupole field and are effectively removed. Iron does not react with the ammonia and continues through the quadrupole to the detector.

The KED uses collisions between an inert gas, like helium, and interfering ions. As the pressure of the gas in the cell is increased, the gas will begin to collide with the ions in the ion path. Helium is a small gas and should collide more with interferences that are larger in size than the ion of interest. The collisions are intended to reduce the kinetic energy of the ions. Those ions with lower kinetic energy can then be prevented from reaching the quadrupole by inserting a voltage barrier between the cell and the quadrupole. Because collisions take place with the analytes also, a reduction of the signal for the analytes of interest occurs.

3.5.2. Perkin Elmer NexION ICP-MS 300D Method

 Digested samples are diluted 10 times (Minimum) using 2 % nitric acid prior to analysis  An aliquot of the sample is carried to the plasma of the ICP-MS to be nebulized.  Mist is pulled into plasma by vacuum and dried, vaporized, and elements ionized.  The ions are then filtered according to mass by the instrument.  The instrument detects each element mass and determines the concentrations.  Dilutions are calculated and results are given.

3.6. Chemicals

A variety of chemicals were used throughout the testwork completed in the area of chemical analysis, mineralogical analysis and metallurgical testing. All reagents used were of analytical grade and supplied from appropriate vendors throughout the chemical industry.

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3.7. Metallurgical Methods

In the metallurgical methods section the standard flotation procedure utilized for each test is described. Particle size analysis was completed on each of the five samples received and is also described.

3.7.1. Flotation Procedure

A standard flotation procedure with minor modifications for reagent screening tests was used in all testwork completed on the final and rougher concentrates. All flotation tests were completed using a modified Denver Flotation Machine with a variable speed motor that utilized the addition of forced air. The tests were typically carried out between 950 and 1100 revolutions per minute (RPM) with ~ 5.0 liters per minute (LPM) of air during flotation. Air flow rates were controlled/changed based upon the observed froth structure of the test being performed, which did vary based upon the reagents/conditions of each test. Standard laboratory kinetic flotation tests were performed which consisted of four stages of five minute intervals for a total retention time of 20 minutes. The pull rate for each flotation test was based upon the observation of the mineralization of the froth. Spray water was added to each test to wash the minerals that become trapped on the sides of the flotation cell back into the slurry. All flotation tests were carried out in a standard 2.3L Denver Flotation Cell that was cleaned prior to each test.

The pH and slurry potentials (ORP) of each test were monitored before, during and after flotation and recorded on laboratory flotation sheets. The pH probe used in for the tests was a standard pH/temperature single junction 12 mm diameter probe. The slurry potential was measured with a standard ORP silver/sliver chloride reference electrode.

Chemical reagents were added differently to each test based upon the conditions sought for each flotation test. The Cytec reagents tested in the program were prepared using deionized water in 1 % weight/weight solutions and added using calibrated analytical grade laboratory pipets. Two different frother types were used for each test, a polyglycol and methyl isobutyl carbinol (MIBC), both were made in deionized water at a concentration of 0.8 g/L. The Nalco reagents were shown not be water soluble and therefore were added to each test neat in micro-syringes based on the dosages specified for each test. A single test required the addition of activated carbon, which was of analytical grade and added neat. Some tests required the adjustment of pH based upon the parameters sought, the pH was modified with hydrated (Ca(OH)2 (S)) or gas (SO2 (g)). All reagents used for the testwork are summarized in Table 3.4 with their active chemical substituents based upon the mineral data safety sheet and information provided from the chemical vendors.

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Table 3.4: Chemical Reagents used in Metallurgical Testwork

Manufacture Chemical Name Active Ingredient Water Soluble Use Flotation Reagents Cytec 7261A polyacrylamide X Depressant of metal sulfides Cytec 7262 polyacrylamide X Depressant of metal sulfides Cytec 3894 Dialkyl Thionocarbamate - Selective collector for Cu Cytec 404 Mercaptobenzothiozole X Cu collector for tarnished Cu minerals Nalco TX-15281 Thio-ester - Selective collector for Cu Nalco TX-15155 Thio-ester - Selective collector for Cu Fisher DG-13 Activated Carbon 13 - Flotation reagent collector SNF Sodium Isopropyl Xanthate Xanthate X Sulfide Collector

Frothers Cytec Aero Froth 65 Polyglycol X Flotation Frother Cytec Methyl Isobutyl Carbinol X Flotation Frother pH Modifiers

Fisher Calcium Hydroxide Ca(OH)2 X Base, flotation pH modifier Sulfur Dioxide SO2 (g) - Acid, flotation pH modifier

Tests requiring the use of SO2(g) were those in which pre-aeration methods were utilized in the conditioning stage prior to flotation. Pre-aeration methods involved the use of air or SO2 or the combination of both at low flow rates, < 1.0 LPM. The pre-aeration condition stage lasted for 30 minutes prior to the flotation. Tests in which air and/or SO2 were used for pre-aeration required the adjustment of pH at the conclusion of conditioning stage with hydrated lime.

During flotation testing the flotation concentrates were filtered individually at the conclusion of each flotation stage, 4 in total. At the conclusion of each flotation test the sample was slurry split to produce sample splits for analytical characterization. The tail samples were also filtered; both the concentrates and tails once filtered were dried in an oven overnight at 60oC. Once the samples were dried they were individually ring and puck pulverized then submitted for chemical analysis in LIMS to the analytical laboratory.

3.7.2. Particle Size Characterization – Screen Analysis

A split from each sample was first wet screened at 500 mesh to produce two screen fractions for each sample, +500 and -500 mesh. The screen fractions were then dried overnight in a hot room (140oF). The plus 500 mesh fraction from each sample was passed through a set of sieves at the following sizes; 65 mesh, 100 mesh, 150 mesh, 200 mesh, 270 mesh, 325 mesh, 400 mesh, and 500 mesh. Each plus 500 mesh fraction was ro-tapped in a Humboldt Sieve Shaker for 15.0 minutes (Steinhart 2010). At the conclusion of the 15.0 minutes, the amount of sample that was retained on each screen was weighed and kept separate for future mineralogical characterization. Table 3.5 summarizes the results from particle size analysis on each of the five samples. All testwork particle size data sheets are provided in Appendix A.

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Table 3.5: Summary of Sample Screen Particle Size Analysis

Rougher Final Copper Rougher Rougher Sample ID Scavenger Rougher feed Concentrate Concentrate Tails Concentrate NMS Lot No. 110322-1 110322-2 110322-3 110322-4 110322-5 Mesh Microns µm % Passing 65 208 100.0 99.2 100.0 93.5 89.0 100 147 100.0 97.6 97.7 83.1 76.0 150 104 96.7 94.1 93.6 73.9 66.3 200 74 88.1 86.3 83.8 64.0 56.9 270 53 77.2 75.8 72.9 55.4 49.2 325 43 72.4 71.3 68.3 52.3 46.0 400 38 68.4 67.0 65.5 51.1 43.9 500 25 56.8 56.3 56.1 42.9 38.0 Pan < 25

P80 59 62 67 135 169

3.7.3. Particle Size Analysis – Cyclosizer Analysis

Three tests required the use of a Warman Cyclosizer to analyze the particle size of the material due to fine particle sizes from re-grinding of the sample prior to flotation testwork. Ten to twenty grams of sample was used for analysis by the Cyclosizer and weighed into a tarred 200 mL beaker. After the sample was weighed approximately 100 mL of deionized water and 3 drops of detergent was added to the beaker containing the sample. The sample was then put into a Cole Parmer 750 watt ultrasonicator operated at 75 % for 1.0 minute to break up agglomerates in the sample. The sample was then transferred to the sample bomb of the Cyclosizer and positioned in place for analysis. Three samples were run on the Cyclosizer for 20 minutes at 21o C with a flow of tap water of 206 millimeters (mm). At the conclusion of 20 minutes the samples were collected for each of size fractions on the Cyclosizer, five in total (40 µm, 30 µm, 20 µm, 15 µm, and 11 µm). Each of samples was fileted into tared ceramic and dried then weighed for analysis to determine the particle size of each test. A specific gravity of 2.7 was assumed for the calculation of each particle size. Table 3.6 presents the particle size for each of three tests characterized by Cyclosizer analysis. All Cyclosizer data sheets are provided in Appendix B.

Table 3.6: Summary of Sample Cyclosizer Particle Size Analysis

Regrind Sample Water Run P80 Sample Product o (min) Weight Flow (mm) Temp C Time (µm) Z10-6-1 0 11.8 206 21 20 74 Float Z10-6-2 5 14.8 206 21 20 21 Tails Z10-6-3 15 15.6 206 21 20 16

24

CHAPTER 4 CHARACTERIZATION OF COPPER ORE AND TESTWORK FEED SAMPLES

Analytical and mineralogical characterization results will be presented and discussed in this section.

4.1. Analytical Characterization of Head Samples

Chemical analysis was completed on two head splits prepared from each of the five samples received. The results are presented in Table 4.1; each value represents the average of the two heads submitted for chemical analysis. The penalty elements of interest are As, Bi, Cd, Sb, Se, and Cu. The arsenic content of the samples ranged between 1072 to 19231 ppm As, the rougher concentrate sample had the highest As content and the rougher tails had the lowest As content. The bismuth content of the samples ranged between 14 and 901 ppm, with the rougher concentrate and the final concentrate having the highest Bi values. Sb content ranged between 7 and 76 ppm Sb; Cd ranged between 8 and 304 ppm Cd. Selenium was also seen in all five samples ranging between 7 and 156 ppm Se, with the highest Se content in the rougher and final concentrate. Copper was also present in all samples ranging between 0.062 and 16.6 % Cu, with the final concentrate having the highest Cu value and the rougher tails having the lowest Cu value.

Based on the information provided from chemical analysis it was determined that the five samples received contained sufficient penalty element values to pursue further testwork. The final and rougher concentrates were selected as the obvious potential separation points between the penalty elements and copper values. The remaining three samples were stored if the need arose for further analysis or testwork.

Each of the five samples had assay by size chemical analysis completed on a head to include three different screen size fractions, +270 mesh, 270 x 500 mesh, and -500 mesh. Table 4.2 shows the breakdown of the penalty element distributions for each of the five products in the three screen fractions. The vast majority of the penalty elements occur in the minus 500 mesh fraction for each sample. The final and rougher concentrates have > than 50 % of the all penalty elements occurring in the minus 500 mesh fraction with > 75 % of the total bismuth occurring in the minus 500 mesh fraction as well. The minus 500 fraction also accounted for the majority of the weight in each of the samples. It was important to note the silver distribution in each sample based on the mineralogy that will be discussed in Section 6.2. The majority of the bismuth occurred in a silver sulfide mineral in all composites analyzed.

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Table 4.1: Sample Head Analytical Characterizations

Rougher Rougher Scav. Product Final Concentrate Rougher Feed Rougher Tails Concentrate Concentrate NMS LOT No. 110322-1 110322-2 110322-3 110322-5 110322-4

As ppm 1197 19231 9223 1279 1072 Sb ppm 8 76 53 10 7 Se ppm 124 156 49 7 7 Au1/ ppm 81.2 35.8 3.19 0.493 0.2155 Bi ppm 648 901 123 23 14 Carbon Roast 550 % <0.01 <0.01 <0.01 0.100 0.090 Sulfur Roat 550 % 9.72 8.77 2.470 0.730 0.560 Carbon Total % 0.09 0.055 0.165 0.25 0.21 Sulfur Total % 24.6 34.7 28.3 5.47 3.59 CAI3/ % 0.07 0.05 0.05 0.03 0.04 CuAS ppm 1528 1984 16 129 122 CuCN3 ppm 12681 16792 1709 419 194 Hg ppm 0.8335 1.47 0.434 0.138 0.049 Ag ppm 533 532 58.55 11.25 5.75 Al ppm 6596 11266 21366 37680 39609 Be ppm <2 <2 <2 <2 <2 Ca ppm 7817 6231 17519 27832 27676 Cd ppm 153 304 44 12 8 Co ppm 37 137 162 30 24 Cr ppm <2 <2 25 63.5 57 Cu % 16.6 14.2 1.02 0.171 0.062 Fe ppm 20.7 31.2 29.9 8.50 6.15 K ppm 3267 6306 12865 30047 30610 Mg ppm 51677 8426 23529 31347 29714 Mn ppm 349 286 444 676 635 Mo ppm 33.5 23 39.5 7 7 Na ppm 243 179 413 1041 1038 Ni ppm 115 319 312 55 49 Pb % 0.227 0.450 0.055 0.012 0.007 Sr ppm 5 21 36 50 54 Ti ppm 372 517 1005 2053 1953 Tl ppm <20 <20 <20 <20 <20 V ppm 34 51 85 131 125 Zn ppm 1.16 3.11 0.446 0.074 0.051 SCIS2/ % 22.1 35.0 28.2 4.34 2.78 So % 0.130 0.065 0.040 <0.01 <0.01

1/ Au by f ire assay with an AA f inish.

2/ SCIS is S= Sulf ur, measured as sulf ur not solubilized by a hot sodium digestion.

3/ Carbon not soluble in dilute HCl. Ref erred to as "organic carbon."

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Table 4.2: Assay-by-Size Head Analytical Characterizations

Assays Distribution 270 270 x 500 -500 Total 270 270 x 500 -500 Total

wt % 22.8 20.4 56.8 100 As ppm 1422 1997 1056 1331 24.4 30.6 45.1 100.0 Sb ppm 1 1 17 10 2.3 2.0 95.7 100.0 Se ppm 131 122 153 142 21.1 17.6 61.3 100.0 Bi ppm 259 448 888 655 9.0 14.0 77.0 100.0 Ag ppm 466 491 628 563 18.9 17.8 63.4 100.0

LOT LOT 110322-1 Cd ppm 112 145 206 172 14.8 17.2 68.0 100.0 Final Concentrate Final Concentrate Cu % 20.336 18.504 16.3736 17.7 26.2 21.3 52.5 100.0

wt % 24.2 19.5 56.3 100 As ppm 8876 23438 23767 20099 10.7 22.7 66.6 100.0 Sb ppm 18 59 94 69 6.3 16.7 76.9 100.0 Se ppm 136 172 210 185 17.8 18.2 64.0 100.0 Bi ppm 327 680 1209 893 8.9 14.9 76.3 100.0 Ag ppm 306 441 654 528 14.0 16.3 69.7 100.0 LOT LOT 110322-2 Cd ppm 276 300 330 311 21.5 18.8 59.7 100.0

Rougher Concentrate Concentrate Rougher Cu % 12.8 14.4 12.8 13.1 23.7 21.4 54.9 100.0

wt % 27.1 16.8 56.1 100 As ppm 5498 15116 9982 9629 15.5 26.4 58.2 100.0 Sb ppm 41 57 47 47 23.6 20.4 56.0 100.0 Se ppm 85 87 63 73 31.6 20.0 48.4 100.0 Bi ppm 125 118 126 124 27.2 15.9 56.8 100.0

110322-3 Ag ppm 62.4 53.6 57.3 58 29.1 15.5 55.4 100.0

Rougher Scav. Rougher Cd ppm 57 58 46 51 30.3 19.1 50.6 100.0 Concentrate LOT Concentrate Cu % 0.910 0.660 0.577 0.681 36.2 16.3 47.5 100.0

wt % 44.6 12.5 42.9 100 As ppm 820 2742 1555 1376 26.6 24.9 48.5 100.0 Sb ppm 7 14 13 10 29.9 16.7 53.4 100.0 Se ppm 11 26 19 16 30.1 19.9 50.0 100.0 Bi ppm 11 37 33 24 21.2 19.3 59.5 100.0

110322-5 Ag ppm 6.6 13.8 16.3 12 25.2 14.8 60.0 100.0

Cd ppm 10 20 19 15 29.5 16.5 53.9 100.0 Rougher Feed LOT Feed Rougher Cu % 0.140 0.261 0.234 0.196 32.0 16.7 51.4 100.0

wt % 50.8 11.2 38 100 As ppm 770 2078 1307 1121 34.9 20.8 44.3 100.0 Sb ppm 7 14 10 9 39.8 17.6 42.6 100.0 Se ppm 13 13 12 13 52.3 11.5 36.1 100.0 Bi ppm 8 19 18 13 31.4 16.4 52.2 100.0

110322-4 Ag ppm 4.1 5.3 7.8 6 36.9 10.5 52.5 100.0 Cd ppm 7 12 13 10 36.1 13.7 50.2 100.0 Rougher TailsRougher LOT Cu % 0.063 0.050 0.076 0.066 48.1 8.5 43.4 100.0

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4.2. Mineralogical Characterization by XRD

Each of the five samples was submitted for mineralogical analysis by XRD and the results are presented in Table 4.3. The major copper mineral identified was chalcopyrite which ranged between 0.5 and 63.6 wt. %. The majority of the chalcopyrite was identified in the final concentrate. Pyrite was present in all samples ranging between 7 and 12 wt. %. The main reason for XRD analysis was for the identification of gangue minerals to assist building the mineral lists for the SEM/MLA runs. Significant talc was also identified in the final concentrate. Often talc is the reason for recovery in the final concentrate due to the association with talc. Fluorine is another penalty element that could be considered for separation testwork but due to the complication of chemical analysis (time-consuming and costly assay) and the association of fluorine with naturally floatable gangue (compared to sulfide association of the suite of penalty elements under consideration) fluorine was not tracked in this study.

Table 4.3: Results of XRD Analysis

Rougher Rougher Final Rougher Rougher Minerals Scavenger Feed Concentrate Concentrate Tails Concentrate Andradite % Biotite % 6.7 1.0 3.0 7.3 Chalcopyrite % 0.5 55.7 42.9 1.3 0.1 Clinochlore IIb % 7.2 7.7 5.2 Diopside % 4.9 6.7 Galena % 0.6 Hornblende % 6.2 3.8 7.0 Muscovite 2M % 3.3 5.4 Orthoclase % 16.5 8.4 16.2 Pyrite % 6.3 6.0 32.2 33.9 4.3 Pyrrhotite % 6.1 12.9 Quartz % 48.5 7.9 10.7 24.6 47.8 % 1.6 6.3 1.2 Talc % 29.0 0.3 3.3 Totals 100 100 100 100 100

Mineral Elemental Composition

Andradite Ca3Fe2Si3O12 Sphalerite ZnS

Biotite K(Mg,Fe)3AlSi3O10(F,OH)2 Magnesiohornblende (Ca,Na)2–3(Mg,Fe,Al)5(Al,Si)8O22(OH,F)2

Chalcopyrite CuFeS2 Muscovite (KF)2(Al2O3)3(SiO2)6(H2O)

Clinochlore IIb Mg 5Al)(AlSi3)O10(OH)8 Orthoclase KAlSi3O8

Diopside MgCaSi2O6 Pyrite Fe2S

Quartz SiO4 Pyrrhotite Fe7S8 Galena Talc PbS Mg 3Si4O10(OH)2

4.3. Mineralogical Characterization by MLA

Each of the five samples was submitted for Mineral Liberation Analysis (MLA) by scanning electron microscope (SEM). The main focus of the MLA discussion will be on the final and rougher concentrates as they were used for metallurgical testwork. The first objective of MLA was to determine the occurrence of the bismuth minerals in the final and rougher concentrates. Once the bismuth minerals were located liberation of the bismuth minerals was of importance. The greater the liberation of the bismuth minerals the greater chance of successfully separating the bismuth minerals from copper minerals. Finally, mineral

28 association was also important due to the potential association of the bismuth minerals with copper minerals. The next few sections will describe in detail the results from MLA analysis of the final and rougher concentrate samples that were used for the metallurgical testwork, but analysis of the remaining circuit samples will also be addressed in Section 4.3.3 Modal Mineralogy.

4.3.1. MLA Calculated Assay vs. Chemical Analysis

It is important to compare the chemical analysis results to the calculated results from MLA. This was done to ensure that the there was good agreement between the analytical results and the MLA results. Table 4.4 compares the MLA results for the final concentrate and Table 4.5 compares the MLA results for the rougher concentrate to those from chemical analysis.

Table 4.4: MLA Calculated Assay vs. Chemical Analysis for Final Concentrate

As (ppm) Bi (ppm) Cd (ppm) Cu (%) Fraction Percent Percent Percent Percent Chem MLA Difference Chem MLA Difference Chem MLA Difference Chem MLA Difference Head 1186 905 26.9 633 959 41.0 185 176 5.0 16.6 19.2 14.4 +270 1422 940 40.8 260 134 64.0 112 52 73.2 20.3 20.4 0.3 270 x 500 1997 1116 56.6 448 908 67.8 145 122 17.2 18.5 19.7 6.3 Final Concentrate -500 1056 817 25.5 888 1309 38.3 206 246 17.7 16.4 18.4 11.7

The final concentrate elemental MLA values presented in Table 4.5 are similar to those obtained from the chemical analysis. With only a few exceptions, mainly in high arsenic values, a good amount of confidence can be placed on the MLA results. Typically MLA and analytical assays for this property are within ~ 20 %, but penalty elements can be as high as 40 % due to their relatively low concentrations.

Table 4.5: MLA Calculated Assay vs. Chemical Analysis for Final Concentrate

As (ppm) Bi (ppm) Cd (ppm) Cu (%) Percent Percent Percent Percent Fraction Chem MLA Difference Chem MLA Difference Chem MLA Difference Chem MLA Difference Head 19231 20032 4.1 902 2031 77.0 304 495 47.8 14.2 14.0 1.1

+270 8876 7823 12.6 888 327 92.3 276 235 16.0 12.8 11.4 11.6 Rougher Rougher

Concentrate 270 x 500 23438 27739 16.8 680 1220 56.8 300 376 22.5 14.4 13.6 5.7 -500 23767 22512 5.4 1210 3037 86.0 330 649 65.2 12.8 15.2 17.1

Rougher concentrate comparisons of MLA to chemical analysis are similar with the exception of the bismuth values. This again is most likely due to the low concentration of bismuth and the relatively small particle size of the minerals that contain bismuth. MLA tends to overestimate the concentration of elements associated with small grain sizes due to the resolution limit of the beam. The property from which the samples were taken has a well-established mineral list based on many years of monthly composite MLA testwork and MLA results for this testwork were not seen as unreliable.

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4.3.2. Distribution of Penalty Element Minerals – Final and Rougher Concentrates

In order to apply successful flotation separation methods it was important to know the distribution of the penalty elements in the samples and the occurrence in minerals. Table 4.6 presents the distribution of penalty elements and their mineral associations in the final concentrate and rougher concentrate.

Table 4.6: Distribution of Penalty Elements in the Final and Rougher Concentrates

Final Concentrate Rougher Concentrate Distribution Distribution Mineral Mineral As (%) Bi (%) Cd (%) Cu (%) Sb (%) As (%) Bi (%) Cd (%) Cu (%) Sb (%) Arsenopyrite 74.88 Arsenopyrite 96.73 Bismuthinite 9.07 Bismuthinite 9.38 Chalcocite Chalcocite 0.01 Chalcopyrite 94.80 Chalcopyrite 91.53 Covellite 4.84 Covellite 7.22 Crandallite Crandallite Enargite 25.12 0.30 Enargite 3.27 1.19 Fe Oxide 0.06 Fe Oxide 0.03 Hessite Hessite Pyrargyrite 0.26 Pyrargyrite 0.30 Sphalerite 99.97 Sphalerite 100.00 Stromeyerite Stromeyerite 0.01 Tetradymite 8.00 Tetradymite 2.76

AgPbBiS3 56.26 AgPbBiS3 82.24 PbBiSbAgS 26.67 0.01 99.74 PbBiSbAgS 5.62 99.70 Total 100.00 100.00 100.00 100.00 100.00 Total 100.00 100.00 100.00 100.00 100.00

In both products, the main copper mineral is chalcopyrite and the main bismuth mineral AgPbBiS3. The majority of the antimony occurs in a lead-bismuth sulfide mineral, PbBiSbAg in both the final and rougher concentrates. The final and rougher concentrates have the same minerals present, but at slightly different distributions compared to one another. Four bismuth minerals were identified in both samples were bismuthinite, tetradymite, AgPbBiS3, and PbBiSbAgS. The vast majority of the Arsenic was shown to be in arsenopyrite and enargite, but arsenic varied between the two minerals in the different samples. The final concentrate contained more enargite than the rougher concentrate which had the majority of the arsenic as arsenopyrite. Both samples also had sphalerite as the main carrier of cadmium and a significant portion of the bismuth and antimony occur in the same mineral, PbBiSbAgS.

4.3.3. Modal Mineralogy

Modal analysis of all five samples was obtained from the MLA to convert each mineral into a volume percentage. Table 4.7 presents the modal mineralogy for each of the five products with the minerals of interest highlighted, complete modal mineralogy for each sample is provided in Appendix. The minerals of interest in each of samples contained the penalty elements that are to be tracked for the separation of copper and bismuth. Table 4.7 focuses on the arsenic and bismuth minerals that were able to be measured in each of the five samples. It was important to show where the penalty elements concentrated based on the flowsheet of the operation. Based on the modal mineralogy and chemical analysis performed on each of the composites the majority of the arsenic and bismuth concentrated in the final

30 tailings, but due to the low weight pull of the final concentrate the overall concentrations of penalty elements in the final copper concentrate still pose a significant problem for the operation.

Table 4.7: Modal Mineralogy of Five Copper Concentrator Circuit Samples

Rougher Rougher Scav. Product Rougher Feed Final Concentrate Rougher Tails Concentrate Concentrate NMS LOT No. 110322-5 110322-1 110322-2 110322-3 110322-4

Amphibole wt. % 6.52 1.74 0.54 2.37 6.13 Andradite wt. % 0.58 0.22 0.07 0.24 0.44 Ankerite wt. % 0.01 Apatite wt. % 0.47 0.13 0.06 0.17 0.35 Arsenopyrite wt. % 0.29 0.15 4.21 2.00 0.23 Barite wt. % 0.01 Biotite wt. % 3.81 0.79 0.46 1.77 3.73 Bismuthinite wt. % 0.01 0.02 Calcite wt. % 0.73 0.02 0.08 0.32 0.61 Chalcopyrite wt. % 0.61 52.48 37.04 2.13 0.22 Chlorite wt. % 0.85 0.47 0.21 0.42 0.86 Covellite wt. % 0.04 1.39 1.52 0.06 0.04 Crandallite wt. % 0.02 0.02 0.04 0.04 Diopside wt. % 7.67 1.69 1.11 3.55 7.88 Electrum wt. % 0.14 Enargite wt. % 0.01 0.12 0.34 0.12 Epidote wt. % 1.12 0.15 0.15 0.43 1.18 Fe Oxide wt. % 0.83 0.91 0.37 0.85 0.84 Forsterite wt. % 0.04 0.14 0.03 0.02 Galena wt. % 0 0.04 0.24 0.01 Illite wt. % 6.94 0.92 1.49 3.62 9.06 Iron wt. % 0.03 0.01 0.16 0.01 Jarosite wt. % 0.05 0.01 0.03 0.02 0.06 K-feldspar wt. % 10.27 0.41 0.83 2.41 10.24 Kaolinite wt. % 0.11 0.02 0.05 0.04 0.19 Lipscombite wt. % 0.01 Mn Oxide wt. % 0.01 Monazite wt. % 0.04 0.01 0.02 0.01 Plagioclase wt. % 1.01 0.05 0.07 0.36 1.1 Plumbogummite wt. % 0.02 Pyrite wt. % 11.54 11.02 35.07 49.37 6.03 Pyrrhotite wt. % 1.34 1.82 6.6 13.37 1.1 Quartz wt. % 42.6 3.41 3.5 11.95 47.64 wt. % 0.9 0.49 0.16 0.44 0.71 Rutile wt. % 0.24 0.07 0.05 0.11 0.26 Smectite wt. % 0.16 0.04 0.04 0.16 0.24 Sphalerite wt. % 0.13 1.53 4.27 0.75 0.05 Stromeyerite wt. % 0.01 Talc wt. % 0.5 19.3 0.72 2.47 0.23 Tetradymite wt. % 0.01 0.01 Titanite wt. % 0.49 0.04 0.05 0.19 0.45 Trolleite wt. % 0.01 0.01 0.03

AgPbBiS3 wt. % 0.02 0.16 0.5 0.04 0.002 PbBiSbAgS wt. % 0.09 0.04 0.002 Zircon wt. % 0.02 0.01 0.01 Total wt. % 100 100 100 100 100

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Table 4.8 presents a detailed modal mineralogy and elemental distribution of the elements of interest that contain the penalty elements as well as the copper distributions in the final concentrate. The final concentrate showed copper minerals to be mainly chalcopyrite and covellite with some enargite. The chalcopyrite content of the final concentrate accounted for 94.8 % of the copper while covellite accounted for 4.8 % of the total copper present in the sample. The modal mineralogy confirmed that the main separation of copper from bismuth minerals had to occur with either the depression or activation of chalcopyrite since it was shown to be the main copper mineral in the samples. Bismuth mineralogy was divided between four different minerals; bismuthinite, tetradymite, AgPbBiS3, and PbBiSbAgS. The main contributors of bismuth content in the final concentrate were AgPbBiS3 accounting for 56.3 % of the total bismuth content and PbBiSbAgS that accounted for 26.7 % of the total bismuth present in the final concentrate.

Table 4.9 presents the detailed modal mineralogy with elemental distributions for the rougher concentrate and showed similar results to the final concentrate. Copper was distributed between four minerals; chalcocite 0.01 %, chalcopyrite 91.5 %, covellite 7.2 %, and enargite 1.19 % of the total copper. Main differences in the copper mineralogy between the final and rougher concentrates were the elemental distribution of copper with enargite (more enargite in the rougher concentrate).

Bismuth mineralogy was also similar to the final concentrate with bismuth being distributed between four minerals; bismuthinite, tetradymite, AgPbBiS3, and PbBiSbAgS. The difference in the rougher concentrate was that the majority of the bismuth was associated with AgPbBiS3, which accounted for 83 % of the total bismuth in the sample. This suggested that in the current plant configuration that a portion of the AgPbBiS3 mineral gets depressed during cleaner flotation. Both the final concentrate and rougher concentrate exhibited similar mineralogical distributions of copper and bismuth minerals with only minor differences. However, as will be presented in the metallurgical results each sample responded differently to flotation conditions.

Arsenic mineralogy was also similar between the final and rougher concentrates among the distributions of arsenopyrite and enargite. The arsenopyrite content in the final concentrate slightly decreased compared to the content shown to be in the rougher concentrate. Current operational conditions at the copper concentrator require an elevated pH in the cleaner flotation stage. The elevated pH and the decrease in the arsenopyrite content suggest that some arsenopyrite is depressed during cleaner flotation, which has also been discussed in Chapter 2, Literature Review. The enargite content was shown to increase in the final concentrate versus the rougher concentrate, showing that enargite is not depressed during cleaner flotation as was shown to be the case with the arsenopyrite content in the copper products.

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Table 4.8: Modal Mineralogy and Elemental Distribution of Final Concentrate

Elemental Distribution Mineral Wt% Area% Particle Count Grain Count Bi (%) Cu (%)

Alunite 0.00 0.00 1 1 Amphibole 1.74 2.13 1364 1841 Andradite 0.22 0.22 152 180 Ankerite 0.00 0.00 8 13 Apatite 0.13 0.16 118 138 Arsenopyrite 0.15 0.09 68 93 Biotite 0.79 0.98 579 719 Bismuthinite 0.01 0.01 11 14 9.07 Calcite 0.02 0.03 77 126 Chalcocite 0.00 0.00 1 1 Chalcopyrite 52.48 47.16 17100 19394 94.80 Chlorite 0.47 0.58 369 411 Covellite 1.39 1.11 875 2012 4.84 Crandallite 0.00 0.01 8 10 Diopside 1.69 1.88 827 1113 0.00 0.00 5 5 Electrum 0.14 0.03 6 6 Enargite 0.12 0.10 102 157 0.30 Epidote 0.15 0.17 152 193 FeOx 0.36 0.32 302 410 FeOx_FeSO4 0.55 0.54 442 585 0.06 Forsterite 0.14 0.16 46 61 Galena 0.04 0.02 13 13 Greenockite 0.00 0.00 1 1 Hessite 0.00 0.00 1 1 Illite 0.84 1.22 1216 1703 Il_Alt FeSulfide 0.08 0.08 157 369 Iron 0.01 0.00 4 4 Jarosite 0.01 0.02 43 45 K-feldspar 0.41 0.63 420 525 Kaolinite 0.02 0.02 19 24 MnOx 0.00 0.00 2 3 Monazite 0.01 0.01 15 24 Plagioclase 0.05 0.07 81 110 Plumbogummite 0.00 0.00 2 2 Pyrargyrite 0.00 0.00 3 3 Pyrite 11.02 8.58 3865 4433 Pyrrhotite 1.82 1.48 730 852 Quartz 3.41 4.82 1909 2361 Siderite 0.49 0.48 380 438 Rutile 0.07 0.06 56 65 Smectite 0.04 0.08 151 202 Sphalerite 1.53 1.37 850 1004 Stromeyerite 0.00 0.00 1 1 Talc 19.30 25.19 8232 8463 Tetradymite 0.01 0.01 13 13 8.00 Titanite 0.04 0.05 117 159

AgPbBiS3 0.16 0.08 163 209 56.26 PbBiSbAgS 0.09 0.05 23 24 26.67 0.01 Zircon 0.00 0.00 8 8

Total 100.00 100.00 31446 48542 100.00 100.00

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Table 4.9: Modal Mineralogy and Elemental Distribution of Rougher Concentrate Elemental Distribution Mineral Wt% Area% Particle Count Grain Count Bi (%) Cu (%)

Amphibole 0.54 0.77 484 632 Andradite 0.07 0.08 59 71 Ankerite 0.00 0.01 6 8 Apatite 0.06 0.08 117 154 Arsenopyrite 4.21 3.02 1230 1421 Barite 0.00 0.00 2 2 Biotite 0.46 0.67 415 491 Bismuthinite 0.02 0.02 20 38 9.38 Calcite 0.08 0.13 82 90 Chalcocite 0.00 0.00 7 10 0.01 Chalcopyrite 37.04 38.25 13015 14672 91.53 Chlorite 0.21 0.30 212 239 Covellite 1.52 1.40 1131 2040 7.22 Crandallite 0.02 0.04 22 28 Diopside 1.11 1.45 586 700 Electrum 0.00 0.00 2 2 Enargite 0.34 0.34 260 351 1.19 Epidote 0.15 0.20 122 183 FeOx 0.17 0.17 162 213 FeOx_FeSO4 0.20 0.22 263 339 0.03 Galena 0.24 0.14 50 57 Hessite 0.00 0.00 3 4 Illite 1.44 2.32 1834 2451 Il_Alt FeSulfide 0.05 0.05 59 63 Iron 0.00 0.00 2 2 Jarosite 0.03 0.04 38 40 K-feldspar 0.83 1.39 680 960 Kaolinite 0.05 0.09 61 71 Monazite 0.02 0.01 25 33 Plagioclase 0.07 0.11 110 170 Plumbogummite 0.02 0.02 5 5 Pyrargyrite 0.00 0.00 1 1 Pyrite 35.07 30.43 10515 11393 Pyrrhotite 6.60 6.20 2406 2646 Quartz 3.50 5.72 2382 3081 Siderite 0.16 0.18 136 158 Rutile 0.05 0.05 117 149 Silver 0.00 0.00 1 1 Smectite 0.04 0.11 177 210 Sphalerite 4.27 4.51 2035 2390 Stromeyerite 0.01 0.00 15 16 Talc 0.72 1.10 439 485 Tetradymite 0.01 0.01 18 23 2.76 Titanite 0.05 0.06 124 224 Trolleite 0.00 0.00 2 2

AgPbBiS3 0.50 0.30 355 476 82.24 PbBiSbAgS 0.04 0.03 32 34 5.62 Zircon 0.00 0.00 6 6

Total 100.00 100.00 30442 46835 100.00 100.00

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4.3.4. Mineral Liberation by Free Surface – Bismuth Minerals

As the bismuth separation technique being investigated was flotation, the liberation of the minerals of interest is of the utmost importance. The following section will describe the liberation of the bismuth minerals in the final and rougher concentrates samples. Although copper liberation by free surface will not be discussed, full liberation by free surface data table for copper are provided in Appendix C.

Liberation data for the four bismuth minerals observed to be in the final concentrate at various stages of liberation are presented below. The tables provide summaries of the free surface of each mineral as well as the distribution of each mineral in a specified liberation class. A liberation class corresponds to the percentage of a specific mineral outer edge that is exposed to the epoxy resin. The outer edge of the mineral exposed to the resin can calculated as a percentage of the entire mineral, those minerals that are > 40% exposed to the resin are assumed to be responsive for recovery in flotation, assuming the surface chemistry is amenable to flotation. So for example, liberation classes for the bismuth minerals ranged from 0 % to 100 % exposed, which 100 % exposed is considered to be fully liberated. In general the majority of the bismuth minerals were greater than 40 % exposed, which suggests the minerals are liberated and should be recoverable in flotation, assuming the appropriate surface chemistry conditions can be created.

Table 4.10 shows the liberation by free surface for the bismuth mineral characterized as PbBiSbAgS, shown to have a total of 23 particles that contained or partially contained the bismuth mineral. Seven particles were shown to have 100 % of their surface free while 4 particles were shown not be exposed (locked). Approximately 96 % of PbBiSbAgS had surfaces > 40.0 % free, meaning the majority of the mineral was liberated and should be amenable to a flotation separation if surfaces are appropriately hydrophobic (flotation) or hydrophilic (depression). However, the average grain size of PbBiSbAgS was ~12 microns which could pose difficulties in a flotation separation.

Table 4.10: Liberation by Free Surface for PbBiSbAgS in the Final Concentrate

PbBiSbAgS Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80= 12 No. of Particles 31423 4 6 1 2 1 2 7 Mean Phase ECD (µm) 0 6 4 33 6 10 11 8 Mean Phase Max Span (µm) 0 9 5 45 9 23 19 11 Particle Distribution (%) 99.8 0.033 0.017 0.004 0.013 0.016 0.034 0.042 Mean Density 3.8 1.9 5.4 1.2 3.1 4.0 4.0 5.4 Distribution of Mineral (%) 0 0.42 1.35 1.56 9.81 13.41 28.36 45.09 Cum. Distn. of Mineral (%) 0 100.0 99.6 98.2 96.7 86.9 73.5 45.1

Figure 4.1 shows a photomicrograph of a liberated PbBiSbAgS particle against a scale of 50 microns. The relatively small particle size of the bismuth containing mineral can be seen based upon the comparison to the other minerals shown in the photomicrograph.

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PbBiSbAgS

Figure 4.1: Photomicrograph of liberated PbBiSbAgS in the final concentrate

Table 4.11 shows the liberated by free surface results for the bismuth mineral characterized as

AgPbBiS3, shown to have a total of 168 particles that contained or partially contained the bismuth mineral. Twenty-six particles were shown to have 100 % of their surface free while 64 particles were shown not be exposed (locked). The unexposed mineral grains were shown to be extremely small, ~3-5µm, inferring that they occur as small inclusions in other minerals. Approximately 70 % of AgPbBiS3 had surfaces > 40.0 % free, meaning the majority of the mineral was liberated and should be amenable to a flotation separation assuming the surface chemistry can be created. AgPbBiS3 was shown to have the largest grain size compared to the three other bismuth minerals, 29 microns. The majority of the bismuth was shown to be associated with AgPbBiS3 in the final concentrate.

Table 4.11: Liberation by Free Surface for AgPbBiS3 in the Final Concentrate

AgPbBiS3 Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80= 29 No. of Particles 31283 64 54 13 2 3 1 26 Mean Phase ECD (µm) 0 3 5 9 17 6 35 9 Mean Phase Max Span (µm) 0 5 7 15 22 9 60 12 Particle Distribution (%) 99.0 0.524 0.316 0.056 0.006 0.012 0.005 0.096 Mean Density 3.75 4.71 4.34 5.75 1.15 4.42 1.40 7.00 Distribution of Mineral (%) 0 2.78 14.0 12.4 2.41 5.08 3.35 59.9 Cum. Distn. of Mineral (%) 0 100.0 97.2 83.2 70.8 68.4 63.3 59.9

Figure 4.2 illustrates a photomicrograph of a partially liberated AgPbBiS3 grain and the relative size of the bismuth mineral compared to other minerals in the sample. Although not illustrated in this photomicrograph, the AgPbBiS3 minerals tend to be coarser when compared to the other three bismuth containing minerals.

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Bis

(AgPbBiS3)

Py

Figure 4.2: Photomicrograph of a partially liberated AgPbBiS3 particle in the final concentrate

Tetradymite (Bi2Te2S) accounted for the least amount of bismuth in the final concentrate. Table 4.12 summarizes the liberation by free surface data showing that only one mineral grain is 100 % liberated, but corresponded to 48 % of the mineral in the sample. The majority of the mineral grains are less than 20 % exposed, with seven of the thirteen particles shown to be locked with other minerals. Mineral association data showed that tetradymite to be associated mainly with talc with minor amounts associated with pyrite and chalcopyrite.

Table 4.12: Liberation by Free Surface for Tetradymite in the Final Concentrate

Tetradymite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80= 8 No. of Particles 31433 7 5 0 0 0 0 1 Mean Phase ECD (µm) 0 3 6 0 0 0 0 7 Mean Phase Max Span (µm) 0 4 9 0 0 0 0 12 Particle Distribution (%) 100 0 0 0 0 0 0 0.006 Mean Density 3.8 1.94 3.68 0 0 0 0 4.14 Distribution of Mineral (%) 0 1.05 50.1 0 0 0 0 48.9 Cum. Distn. of Mineral (%) 0 100.0 99.0 48.9 48.9 48.9 48.9 48.9

Bismuthinite (Bi2S3) contained the second lowest amount of bismuth in the final concentrate and Table 4.13 summarizes the liberation by free surface data. The majority of the particles scanned were 100% liberated by free surface, but only accounted for two particles. The majority of the bismuthinite particles were less than 20 % liberated by free surface. Mineral grain size for bismuthinite minerals was ~16 microns which was smaller than the mineral grain size shown for the main two bismuth minerals, but slightly larger than the tetradymite mineral grain size. It is also important to point out the fine mineral grain sizes of the 100% liberated material, shown to be an average of ~2 microns.

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Table 4.13: Liberation by Free Surface for Bismuthinite in the Final Concentrate

Bismuthanite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80= 16 No. of Particles 31435 3 5 0 1 0 0 2 Mean Phase ECD (µm) 0 5 6 0 16 0 0 5 Mean Phase Max Span (µm) 0 7 9 0 28 0 0 7 Particle Distribution (%) 100 0 0 0 0.002 0 0 0.006 Mean Density 3.8 2.12 5.14 0 1 0 0 3.71 Distribution of Mineral (%) 0 3.06 31.0 0 9.87 0 0 56.0 Cum. Distn. of Mineral (%) 0 100.0 96.9 65.9 65.9 56.0 56.0 56.0

Figure 4.3 illustrates a bismuthinite mineral associated with pyrite; figure 4.3 accurately shows the relative small size of the mineral compared to other minerals present in the sample.

Bis Py

Figure 4.3: Photomicrograph of a partially liberated Bismuthinite particle in the final concentrate

A significant amount of the mineralogical testwork was completed utilizing the rougher concentrate. Table 4.14 shows the liberation by free surface data for PbBiSbAgS. Similar to the mineral found in the final concentrate, PbBiSbAgS had a mineral grain size of ~12 microns. The majority of the mineral was > 40 % liberated (64.4 %). However, a good portion of the mineral scanned was shown not to be exposed, but locked with other minerals. PbBiSbAgS shown from modal mineralogy only accounted for 5.6 % of the total bismuth present in the rougher concentrate.

Table 4.14: Liberation by Free Surface for PbBiSbAgS in the Rougher Concentrate

PbBiSbAgS Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80= 12 No. of Particles 30410 14 8 2 1 7 Mean Phase ECD (µm) 4 4 7 25 8 Mean Phase Max Span (µm) 5 6 14 64 10 Particle Distribution (%) 100 0 0 0 0 0 Mean Density 4.3 3.727 4.350 4.178 1.681 5.320 Distribution of Mineral (%) 3.5 2.6 29.5 1.5 62.9 Cum. Distn. of Mineral (%) 100.00 96.45 93.84 64.37 62.88 62.88 62.88

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The majority of the bismuth present in the rougher concentrate was found to be contained in the mineral AgPbBiS3, over 83 %. Table 4.15 shows the liberation by free surface data. The majority of the particles scanned were shown to be fully liberated or having 100 % of their surface exposed to the epoxy resin. A total of 355 particles were scanned and were shown to have some of their content to be

AgPbBiS3. Approximately 72 % of the particles had > than 40 % of their surfaces exposed, which would be a benefit for a flotation separation. A significant portion of the particles also were < 20 % exposed or fully locked with other minerals, mainly chalcopyrite and pyrite.

Table 4.15: Liberation by Free Surface for AgPbBiS3 in the Rougher Concentrate

AgPbBiS3 Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80= 15 No. of Particles 30087 118 105 30 11 4 1 86 Mean Phase ECD (µm) 4 5 9 8 8 31 8 Mean Phase Max Span (µm) 5 7 13 14 12 35 11 Particle Distribution (%) 95 3 1 0 0 0 0 0 Mean Density 4.3 4.536 4.366 5.225 4.617 3.633 1.398 7.000 Distribution of Mineral (%) 3.9 6.4 17.6 4.1 5.7 0.7 61.6 Cum. Distn. of Mineral (%) 100.00 96.09 89.68 72.08 67.95 62.22 61.56

Figure 4.4 shows the photomicrograph of a partially liberated AgPbBiS3 particle with a decent particle size. The majority of the bismuth in the rougher concentrate occurred in the bismuth mineral AgPbBiS3 making the liberation important for the separation of the mineral. Figure 4.5 is a false color image of the sample particle taken from the MLA. Depicted from the false color image is the association of AgPbBiS3 with pyrite, illite, rutile, and sphalerite.

AgPbBiS3

Figure 4.4: Photomicrograph of a partially liberated AgPbBiS3 particle in the rougher concentrate

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Figure 4.5: False color image of AgPbBiS3 particle

Figure 4.6 shows a fully liberated AgPbBiS3 particle surrounded by particle of chalcopyrite, pyrite and covellite in the rougher concentrate sample.

Gn Ccp Py Bi

AgPbBiS3

Py

Ccp Po Cv

Figure 4.6: Photomicrograph of a liberated AgPbBiS3 particle in the rougher concentrate

As with the tetradymite found in the final concentrate, the number of grains found in the rougher concentrate was low and the majority of grains were not exposed. However, three of the grains that were found were shown to account for ~79 % of the total mineral being 100 % liberated. Table 4.16 shows the liberated by free surface for tetradymite minerals found in the rougher concentrate. The minerals that were shown to be locked were mainly associated with chalcopyrite and pyrite.

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Table 4.16: Liberation by Free Surface for Tetradymite in the Rougher Concentrate

Tetradymite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80= 8 No. of Particles 30424 8 5 2 3 Mean Phase ECD (µm) 4 4 6 5 Mean Phase Max Span (µm) 5 5 9 8 Particle Distribution (%) 100 0 0 0 0 Mean Density 4.3 2.167 2.133 3.927 4.071 Distribution of Mineral (%) 2.2 3.6 14.2 80.0 Cum. Distn. of Mineral (%) 100.00 97.78 94.19 79.96 79.96 79.96 79.96

Bismuthinite that was found to be in the rougher concentrate was also at low concentrates as was the case in the final concentrate. Table 4.17 shows the liberated by free surface for bismuthinite minerals found in the rougher concentrate. In all 20 particles were scanned and shown to contain bismuthinite, 5 of those mineral grains were shown to account for ~65 % of fully liberated bismuthinite minerals. Different for the rougher concentrate, the mineral grain size of the bismuthinite minerals were shown to be coarser at 40 microns.

Table 4.17: Liberation by Free Surface for Bismuthinite in the Rougher Concentrate

Bismuthinite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80= 40 No. of Particles 30422 3 10 1 1 5 Mean Phase ECD (µm) 6 6 4 15 8 Mean Phase Max Span (µm) 8 10 7 19 14 Particle Distribution (%) 100 0 0 0 0 0 Mean Density 4.3 1.958 2.255 2.491 1.366 4.940 Distribution of Mineral (%) 0.9 6.9 23.4 3.1 65.7 Cum. Distn. of Mineral (%) 100.00 99.06 92.14 68.79 65.68 65.68 65.68

4.3.5. Mineral Association – Bismuth Minerals

The objective of the testwork was to separate the bismuth minerals from the copper minerals so it was key to identify if bismuth and copper minerals were associated with each other. Data obtained from the MLA can show the mineral associations based on each product. Mineral association is defined as minerals that are adjacent to each other when the sample was measured on the MLA, which means the minerals occur together in a particle and are not completely liberated from one another. Table 4.18 shows the minerals associations between the copper and bismuth minerals in the final and rougher concentrates. Table 4.18 should be read from left to right based on the mineral specified in the left column, a complete row will account for 100 % of the mineral association within the sample.

The final concentrate was shown to have some mineral associations between copper and bismuth minerals, but at very low levels. The majorities of the copper minerals had free surfaces or were associated with other non-bismuth minerals in the final concentrate. The rougher concentrate generally exhibited the same mineral associations as the final concentrate.

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Both the final and rougher concentrates had the vast majority of the bismuth minerals occurring as free surfaces, meaning no mineral associations. The final concentrate showed the bismuth minerals to be associated with chalcopyrite at low levels with AgPbBiS3 having the highest association at 11.9 wt. %. A good portion of the bismuth minerals were also shown to be associated with gangue minerals and not copper minerals. The rougher concentrate showed similar copper association between chalcopyrite and bismuth minerals, but bismuthinite was shown to have the highest association with chalcopyrite at 15.3 wt. %. Overall, the bismuth and copper mineral associations were low suggesting the liberation of bismuth minerals from copper minerals was not an issue in the final or rougher concentrates.

Table 4.18: Mineral Association in Final and Rougher Concentrate Samples

Final Concentrate

Mineral Chalcocite Chalcopyrite Copper Covellite Bismuthinite AgPbBiS3 PbBiSbAgS Tetradymite Gangue Free Surface Total Chalcocite 100 100 Chalcopyrite 0.76 0.05 7.98 91.20 100 Covellite 0.01 17.14 0.16 10.35 72.34 100 Bismuthinite 3.14 3.02 27.78 66.05 100 AgPbBiS3 11.96 1.28 0.37 6.06 13.58 66.75 100 PbBiSbAgS 1.86 14.06 5.03 79.06 100 Tetradymite 0.58 23.75 75.68 100 Other 7.18 0.46 0.01 0.05 0.01 0.01 92.29 100

Rougher Concentrate Mineral Chalcocite Chalcopyrite Copper Covellite Bismuthinite AgPbBiS3 PbBiSbAgS Tetradymite Gangue Free Surface Total Chalcocite 10.57 22.4 9.36 57.67 100 Chalcopyrite 1.65 0.02 0.11 13.53 84.69 100 Covellite 0.02 22.87 0.09 8.34 68.67 100 Bismuthinite 15.27 5.75 0.04 0.09 5.48 73.38 100 AgPbBiS3 6.76 0.46 0.3 0.46 11.53 80.5 100 PbBiSbAgS 0.67 0.03 5.08 17.69 76.53 100 Tetradymite 5.5 0.1 0.02 7.48 86.9 100 Other 8.52 0.38 0 0.12 0.01 90.96 100

4.3.6. Penalty Element Transport – Flowsheet Distribution

Based on the five products that were sent from site a basic metallurgical balance could be constructed to determine what the deportment of the penalty elements are in a simplified operations flowsheet. Figure 4.7 shows a simplified flowsheet of the operation and a breakdown of where the penalty elements distribution. The flowsheet displays the penalty element and copper distributions for each product. The metallurgical balance was completed based upon flowrate data provided by the operation for the majority of the product streams, with the exception of the individual rougher and scavenger concentrates that had to be calculated based upon the combined rougher concentrate flowrate. The copper was balanced first and the flowrates for all streams were then fixed, the remaining elements were then balanced based upon the new flowrates. Based on Figure 4.7 the majority of the bismuth ends up in the rougher tails with a good portion also concentrating in the final concentrate. The remaining bismuth then concentrates in the final concentrate and cleaner tail products.

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Ro Feed Dist. (%) Final Tails Dist. (%) Cu As Bi Sb Rougher Cu As Bi Sb 100.0 100.0 100.0 100.0 Feed 5.1 80.9 56.2 67.5 Se Cd Pb Se Cd Pb 100.0 100.0 100.0 27.6 56.3 52.4 Rougher Scavenger Final Tails Flotation Flotation

Rougher Scav Con Dist (%) Cu As Bi Sb Rougher Concentrate Scavenger Concentrate 15.1 18.3 13.3 12.4 Rougher Con Dist. (%) Se Cd Pb Cu As Bi Sb 17.8 9.1 11.5 79.8 14.5 37.0 7.3 Combined Se Cd Pb Rougher 21.5 24.4 35.7 Con

Cleaner Tails Dist (%) Cu As Bi Sb Cleaner Cleaner Tails 18.6 18.4 22.4 31.9 Flotation Se Cd Pb 45.0 33.8 33.2

Final Concentrate Final Conc Dist (%) Cu As Bi Sb 76.3 0.7 21.4 0.6 Se Cd Pb 27.4 9.9 14.4

Figure 4.7: Flowsheet diagram of penalty elements

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CHAPTER 5 METALLURGICAL TESTWORK RESULTS

Penalty element separation testwork was completed on two samples, a final and rougher concentrate received from a copper flotation operation. The focus of the testwork was to evaluate flotation conditions that would separate the penalty elements from the final (copper) concentrate due to the ease of implementing a potential solution in operation. As shown in Section 6.1 Analytical Characterization the bismuth concentrated in both the final and rougher concentrates at levels that could be measured using standard analytical methods. A test program was developed and initiated on the final concentrate. Later testing focused on the rougher concentrate to investigate if a better separation between copper and bismuth could be achieved earlier in the process. The subsequent sections will describe the metallurgical approach followed as well as discuss the results obtained from testwork. All testwork flotation data sheets are provided in detail in Appendix D.

5.1. Metallurgical Testwork Approach

Once detailed analytical and mineralogical analysis was completed on the final and rougher concentrates to confirm measurable levels of bismuth were present, metallurgical testwork was initiated. As flotation is the main physical separation processing being utilized to produce a copper concentrate in the operation, the focus of this study was on physical separation processes, i.e. flotation. The subject of a potential hydrometallurgical solution for bismuth removal has been evaluated to some extent previously, and was not considered in this study. The first step was to apply separation methods to the final concentrate and evaluate the metallurgical response. Several baseline flotation tests were completed on the final concentrate without the use of reagent additions, i.e. the final concentrate was refloated without any chemistry or liberation modifications. A baseline grind series was also conducted without the use of reagents to study the impact of mineral surface polishing/conditioning as mineralogical liberation was not seen to be an issue between the copper and bismuth minerals based on detailed mineralogical analysis previously discussed. Once baseline flotation tests were completed reagent trials began to determine if the separation of copper and bismuth minerals could be accomplished.

The literature previously presented described the use of pre-aeration as an important factor in considering a penalty element separation; it was considered an important aspect of the approach followed. Many reagent vendors have also claimed to have made progress in synthesizing reagents targeted/beneficial to penalty element separations and therefore evaluation of reagents was also a significant aspect considered in this study. Numerous tests were completed under a variety of different conditions with a variety of different reagents that were obtained from different vendors. Reagents were sourced based upon conversations with the different chemical vendors and internal Newmont personnel conversations (Gathje 2010). Since the majority of the bismuth was shown to occur in combination with

44 lead, representatives at the chemical manufactures suggested trying lead depression reagents (Thomas 2010). Two reagents were selected from Cytec Industries and another two reagents from Nalco Industries (Laney 2010). All reagents were tested using pre-aeration techniques and pH modifications on both the final and rougher concentrates. Pre-aeration has been shown in the past to aid in surface preparation of the minerals for the addition of reagents (Bulatovic 2007). Two pre-aeration parameters were utilized in the testwork that included the use of air or an air/SO2 combination prior to flotation. The final concentrate was selected as the first sample to initiate testwork on. When the focus shifted to the rougher concentrate the set of parameters had been narrowed down and less screening testwork was needed to find a separation point. Table 5.1 provides brief summaries of all of the tests completed on the final and rougher concentrates and brief comments about the goal of each test.

Table 5.1: Metallurgical Testwork Summary

Final Concentrate Test ID Comments Z10- Baseline Test Series 6-1 Baseline flotation test, no reagents. Natural pH 6-2 Baseline flotation test, no reagents. Slight regrind to polish mineral surfaces. Natural pH. 6-3 Baseline flotation test, no reagents. 15.0 minute regrind , natural pH. 8-13 Repeat of Z10-6-1 baseline test. Reagent Screening Series 8-1 Slight regrind, NaSH to depress Cu minerals. 8-2 No regrind, activated carbon to remove reagent and NaSH for Cu depression. 8-3 Repeat of Z10-8-2, but with Cytec 404 to float silver containing minerals. (Bismuth association) Pre-Aeration Reagent Series 8-4 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and 7262 for bismuth depression. 8-5 Repeat of Z10-8-4, but with SO2 in place of air in preaeration. 8-6 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and 7261A for bismuth depression. 8-7 Repeat of Z10-8-6, but with SO2 in place of air in preaeration. Pre-Aeration Series 8-16 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and Nalco TX-15281 for bismuth depression. 10-1 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and Nalco TX-15155 for bismuth depression.

Rougher Concentrate Baseline Series 8-12 Baseline flotation test, no reagents. Natural pH 10-2 Repeat of baseline flotation test, no reagents. Natural pH Pre-Aeration Reagent Series 8-8 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and Nalco TX-15155 (selective copper collector). 8-9 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and Nalco TX-15281 (selective copper collector). 8-10 Repeat of Z10-8-9, but with SO2 in place of air in preaeration. 8-11 Repeat of Z10-8-8, but with SO2 in place of air in preaeration. 8-14 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper flotation and 7261A for bismuth depression. 8-15 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper flotation and 7262 for bismuth depression. 10-5 SO2 preaeration with no reagents. 10-6 Repeat of Z10-8-14, but with SO2 in place of air in preaeration. Pre-Aeration Series 10-3 Preaeration with no reagents. 10-4 Repeat of Z10-8-15, but with SO2 in place of air in preaeration.

5.2. Baseline Metallurgical Testwork Results- Final Concentrate

Four baseline flotation tests were completed on the final concentrate to evaluate the response of initial baseline copper and bismuth separations. Two baseline tests were completed without reagents and without a regrind of the final concentrate prior to flotation testing. The remaining two baseline flotation tests involved a slight regrinding of the final concentrate prior to flotation to polish the mineral surfaces, but again no reagents were added during flotation testing. Table 5.2 presents the results from each of the baseline tests completed on the final concentrate.

45

Table 5.2: Baseline Metallurgical Results for Final Concentrate

Test Test Ret. Assay Distribution, % Reject Test Regrind P80 Time Ratio Conditions Z10- Date Product wt. As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn (mins) % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Final Concentrate Assayed Feed 100 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 6-1 No 59 1 1CC1 14.9 209 248 124 16 42 252 12.3 14.9 0.125 0.747 3.1 6.1 9.9 7.2 8.5 7.3 10.6 10.1 7.7 10.1 2 1CC1 - 1CC2 34.7 257 310 149 20 50 312 14.4 16.6 0.150 0.87 8.9 17.7 27.6 21.0 23.5 21.1 28.7 26.2 21.6 26.2 3 1CC1 - 1CC3 59.0 476 433 180 27 62 404 17.6 19.9 0.193 1.05 28.1 42.2 56.8 48.2 49.9 46.5 59.9 53.5 47.3 55.9 BASELINE 5 1CC1 - 1CC4 74.3 740 496 187 30 66 439 18.2 21.6 0.209 1.09 55.1 60.8 74.5 66.7 66.5 63.7 77.6 72.8 64.4 73.2 TEST/ FINAL 5 1CC1 - 1CC5 81.1 846 516 187 30 68 452 18.1 22.1 0.212 1.09 68.8 69.1 81.3 74.8 74.5 71.5 84.4 81.4 71.2 79.9 1.22 CONC 24-Apr-12 16 1CT (Average) 43.8 1647 990 185 44 100 776 14.5 21.5 0.367 1.18 30.9 18.7 25.2 25.5 28.5 15.6 18.6 28.8 20.1 0.0 Calc Head 997 605 187 33 74 513 17.4 22.0 0.241 1.11 6-2 X 21 1 1CC1 7.7 203 181 88 11 48 212 9.23 11.2 0.093 0.574 1.5 2.2 3.6 2.6 3.9 3.1 4.1 3.9 2.8 3.9 5.0 mins 2 1CC1 - 1CC2 19.9 220 213 116 13 51 266 12.6 14.0 0.111 0.717 4.2 6.8 12.4 8.2 10.8 10.2 14.6 12.7 8.7 12.7 BASELINE 3 1CC1 - 1CC3 35.2 267 284 143 18 62 331 16.0 16.9 0.142 0.87 9.0 16.0 27.2 19.5 23.1 22.4 32.9 27.2 19.6 27.0 TEST/ FINAL 5 1CC1 - 1CC4 50.7 368 380 163 24 73 381 17.9 18.9 0.180 0.97 17.8 30.8 44.7 37.1 39.2 37.1 53.0 43.6 35.8 43.2 CONC WITH 6-2 5 1CC1 - 1CC5 58.7 452 444 172 27 79 407 18.4 19.6 0.194 1.02 25.3 41.7 54.7 48.7 49.0 46.0 63.0 52.4 44.8 52.6 1.51

REGRIND 24-Apr-12 16 1CT (Average) 41.3 1910 884 203 41 117 680 15.3 25.4 0.340 1.30 74.7 58.3 45.3 51.3 51.0 54.0 37.0 47.6 55.2 47.4 Calc Head 951 588 129 30 86 498 16.6 21.8 0.237 1.07 6-3 X 16 1 1CC1 6.4 211 83 41 10 19 105 3.3 6.4 0.055 0.228 1.4 0.9 2.0 2.1 1.4 1.3 1.3 1.9 1.5 1.9 15 mins 2 1CC1 - 1CC2 16.1 211 156 69 12 34 169 6.9 9.8 0.076 0.420 3.6 4.3 8.6 6.6 6.4 5.4 6.7 7.3 5.2 7.3 BASELINE 3 1CC1 - 1CC3 28.1 224 210 96 14 43 233 10.4 12.9 0.101 0.58 6.6 10.0 20.9 13.3 14.2 13.1 17.6 16.6 12.0 15.4 TEST/ FINAL 5 1CC1 - 1CC4 42.2 246 285 123 18 55 298 13.9 15.9 0.138 0.74 10.9 20.5 40.3 25.0 27.1 25.2 35.3 30.9 24.5 29.2 CONC WITH 6-3 5 1CC1 - 1CC5 49.4 276 340 135 21 62 332 15.1 17.1 0.158 0.80 14.3 28.6 51.7 33.9 35.7 32.9 45.0 38.9 32.9 37.2 1.57

REGRIND 24-Apr-12 16 1CT (Average) 50.6 1610 830 123 40 110 660 18.1 26.3 0.314 1.33 85.7 71.4 48.3 66.1 64.3 67.1 55.0 61.1 67.1 62.8 Calc Head 951 588 129 30 86 498 16.6 21.8 0.237 1.07 8-13 No 59 5 1CC1 32.6 183 233 56 7 21 194 7.5 9.7 0.107 0.374 6.3 11.5 12.0 9.1 11.5 12.5 14.0 14.2 14.6 14.2 5 1CC1 - 1CC2 47.5 264 340 91 14 34 299 13.1 14.6 0.147 0.614 13.2 24.5 28.6 25.7 27.0 28.1 35.4 30.9 29.3 30.9 BASELINE 5 1CC1 - 1CC3 54.9 319 402 111 17 38 344 14.9 16.2 0.165 0.746 18.4 33.5 40.1 37.2 35.0 37.4 46.5 39.8 38.1 39.2 TEST/ FINAL 8-13 5 1CC1 - 1CC4 61.8 398 459 127 20 41 380 16.4 17.8 0.182 0.848 25.8 43.0 51.9 48.3 42.6 46.5 57.5 49.0 47.1 50.1 1.34

CONC 20 10-Oct-12 1CT (Average) 38.2 1856 980 191 34 90 709 19.5 29.9 0.330 1.37 74.2 57.0 48.1 51.7 57.4 53.5 42.5 51.0 52.9 49.9 Calc Head 955 658 152 25 60 506 17.6 22.4 0.239 1.05

Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Baseline Z10-6-1 1.23 1.22 1.04 1.13 1.13 Baseline Z10-6-2 2.49 1.51 1.15 1.29 1.29 Baseline Z10-6-3 3.15 1.57 0.87 1.33 1.26 Baseline Z10-8-13 2.23 1.34 1.11 1.19 1.35

46

Table 5.2 indicates that if the final concentrate is reground the separation between the bismuth minerals and copper minerals is slightly better. A rejection ratio was calculated for each test by taking the cumulative copper recovery and dividing it by the cumulative bismuth recovery for each test, which is presented to the right in Table 5.2. A larger rejection ratio signals a better separation of the copper from bismuth. The rejection ratios for each test slightly increased for the final concentrates that were reground. Rejection ratios for all other penalty elements are presented at the bottom of Table 5.2. Arsenic, antimony and selenium were observed to behave similar to bismuth in this test series, with the exception of cadmium in Test Z10-6-3 with a 15.0 minute regrind. Table 5.2 shows the assayed feed for the final concentrate and that can be compared to each tests calculated assay for each element. In all baseline tests the elemental assayed values compared well to the calculated values for each test, giving confidence in the data for the metallurgical balance of each test.

Baseline Tests Z10-6-1 and Z10-8-13 utilized the same flotation procedure, but had slightly different flotation retention times, 16 minutes versus 20 minutes. No regrinds were present in either of these baseline tests, but the rejection ratios for bismuth from each were only slightly different, 1.22 versus 1.34. This difference could have been caused due to slight sample degradation over time causing less copper and bismuth to float as Test Z10-8-13 was completed many months after Test Z10-6-2. The rejection ratios for the remaining penalty elements, with the exception of arsenic, are similar to baseline tests without a regrind. Test Z10-6-1 was completed in April of 2012 and Test Z10-8-13 was completed in October of 2012.

Figures 5.1 and 5.2 show the cumulative grade recovery curves for copper and bismuth for the four baseline flotation tests on the final concentrate. The data points in the graphs represent a cumulative value (grade or recovery) of the element being presented based on the timed cumulative flotation concentrates for each flotation test. Generally, each flotation test resulted in a total of four concentrates taken over 20 minute flotation period. The concentrates were assayed individually and were used to calculate a cumulative value for each element based on the cumulative flotation time for each concentrate.

The cumulative grade recovery curves for the baseline flotation tests shown in Figures 5.1 and 5.2 indicate that higher copper recoveries are achieved without a regrind prior to flotation. This seems to suggest that liberation or surface polishing does not pose a problem with the final concentrate for copper flotation. The baseline tests do not follow the characteristic grade recovery relationship typically observed, where an increase in recovery would decrease grade over time. The reason for this could be due to the fact the coarser poorly liberated copper particles float initially and the fine liberated copper minerals begin to float towards the end of each baseline float test.

47

In each baseline test copper collection reagents were not added due to the reagent scheme utilized at the copper concentrator where the samples were taken, it was important to mimic the plant conditions in order to establish accurate baselines. The final concentrate has a relativity fine particle size and fine particles are well-known to respond with much slower flotation kinetics which could account for the response from the baseline tests that had a regrind (Fabiano Capponi 1999). However, the baseline flotation test that was completed in October of 2012 was shown to have a lower bismuth recovery and grade than the baseline test without the regrind that was completed in April of 2012. This is most likely due to the aging of the sample in between baseline tests since flotation separations center on the surface chemistries of the minerals. Aging of the surfaces of the minerals can hinder the performance during flotation by the increased presence of an oxidation layer on the minerals, which prevents the mineral from being recovered during flotation. The baseline flotation test was completed in October to cross-check the results of the original baseline test that was completed in April 2012. When the decision was made to trial additional reagent conditions later in 2012 (October) another Baseline Test, Z10-8-13, was completed when a new series of reagent screening tests were completed.

Cumulative Copper Grade vs. Recovery 20.0

18.0

16.0

14.0

12.0

10.0

8.0

6.0 Cumulative Copper Grade, %

4.0

2.0

0.0 0.0 10.0 20.0 30.0 40.0 50.0 60.0 70.0 80.0 90.0 100.0 Cumulative Copper Recovery, % Baselibne April 2012 Baseline October 2012

Baseline April 2012 5.0 minute regrind Baseline April 2012 15.0 minutes regrind

Figure 5.1: Cumulative copper grade recovery curves for baseline tests on final concentrate

48

Cumulative Bismuth Grade vs. Recovery

600

500

400

300

200 Cumulative Bismuth Grade, %

100

0 0.0 10.0 20.0 30.0 40.0 50.0 60.0 70.0 80.0 90.0 100.0 Cumulative Bismuth Recovery, % Baselibne April 2012 Baseline October 2012

Baseline April 2012 5.0 minute regrind Baseline April 2012 15.0 minutes regrind

Figure 5.2: Cumulative bismuth grade recovery curves for baseline tests on final concentrate

49

5.3. Depression of Copper Minerals – Final Concentrate

Three flotation tests were completed on the final concentrate to evaluate different reagent conditions to establish whether the copper bearing minerals could be depressed while the bismuth minerals were selectively floated. Table 5.3 summarizes the three flotation tests that were completed and brief comments of the flotation parameters tested. Baseline Test Z10-6-1 was completed prior to the flotation testwork in this series and was deemed the best baseline test to compare results against

Table 5.3: Summary of Metallurgical Copper Depression Screening Tests for Final Concentrate

Test ID Comments Z10-

Baseline Test Series 6-1 Baseline flotation test, no reagents. Natural pH Reagent Screening Series 8-1 Slight regrind, NaSH to depress Cu minerals. Conditon with nitrogen prior to flotation. 8-2 No regrind, activated carbon to remove reagent and NaSH for Cu depression. 8-3 Repeat of Z10-8-2, but with Cytec 404 to float silver containing minerals. (Bismuth association)

Several flotation parameters were first evaluated to depress copper minerals while attempting to float the bismuth minerals. The first Test, Z10-8-1 utilized the addition of sodium (NaSH) to depress the copper minerals while trying to float the bismuth minerals. Test Z10-8-2 utilized the addition of activated carbon to collect any lingering reagent that was in the sample followed by the addition of NaSH to depress the copper minerals. Test Z10-8-3 was a repeat of Z10-8-2, but utilized the addition of Cytec 404 (mercaptobenzothiozole) to aid in the flotation of the bismuth minerals. Since the majority of the bismuth was contained in a lead silver sulfide mineral, Cytec 404 targeted the flotation of the silver bearing bismuth sulfide (Thomas 2010).

Results for these three tests are presented in Table 5.4. Results summarized in Table 5.4 show that each of the three tests exhibited a decrease in the rejection ratios for the majority of penalty elements with the exception of cadmium compared to baseline flotation test on the final concentrate. The rejection ratios for the three tests in this series were calculated based upon the cleaner tails since the objective was to float the bismuth and leave the copper in the cleaner tails. The rejection ratio calculation was discussed previously in Section 5.2, Baseline Metallurgical Testing on the Final Concentrate. The parameters for the copper depression test series on the final concentrate were shown not to offer a separation that was worth additional scoping tests and no further tests were completed following this separation approach. Although it was important to trial this separation since the bismuth minerals were sufficiently less in quantity compared to the copper minerals present in the final concentrate.

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Table 5.4: Metallurgical Results for Final Concentrate- Copper Mineral Depression Series

Test Test Ret. Assay Distribution, % Reject % Test Conditions Regrind P80 Time Ratio Z10 Date Solids Product wt. As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn (mins) % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Final Concentrate Assayed Feed 100 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 6-1 No 59 1 10.0 1CC1 14.9 209 248 124 16 42 252 12.3 14.9 0.125 0.747 3.1 6.1 9.9 7.2 8.5 7.3 10.6 10.1 7.7 10.1 2 1CC1 - 1CC2 34.7 257 310 149 20 50 312 14.4 16.6 0.150 0.87 8.9 17.7 27.6 21.0 23.5 21.1 28.7 26.2 21.6 26.2 3 1CC1 - 1CC3 59.0 476 433 180 27 62 404 17.6 19.9 0.193 1.05 28.1 42.2 56.8 48.2 49.9 46.5 59.9 53.5 47.3 55.9 BASELINE TEST/ 5 1CC1 - 1CC4 74.3 740 496 187 30 66 439 18.2 21.6 0.209 1.09 55.1 60.8 74.5 66.7 66.5 63.7 77.6 72.8 64.4 73.2 FINAL CONC 5 1CC1 - 1CC5 81.1 846 516 187 30 68 452 18.1 22.1 0.212 1.09 68.8 69.1 81.3 74.8 74.5 71.5 84.4 81.4 71.2 79.9 1.22

24-Apr-12 16 1CT (Average) 18.4 1647 990 185 44 100 776 14 21.5 0.367 1.18 30.9 18.7 25.2 25.5 28.5 15.6 18.6 28.8 20.1 0.0 Calc Head 997 605 187 33 74 513 17 22.0 0.241 1.11 8-1 n/a 59 1 10.0 1CC1 10.7 109 98 11 8 13 44.0 1.26 4.34 0.048 0.152 1.3 1.8 0.7 2.5 1.9 0.9 0.9 2.4 2.2 2.4 2 1CC1 - 1CC2 19.0 107 96 16 8 12 46.6 1.25 4.38 0.046 0.145 2.3 3.1 1.7 4.2 3.1 1.7 1.6 4.3 3.7 4.3 3 1CC1 - 1CC3 23.6 143 110 21 8 15 56.4 1.41 5.53 0.050 0.160 3.8 4.4 2.7 5.7 4.7 2.6 2.2 6.7 5.0 3.7 1. Sodium sulfide to 5 1CC1 - 1CC4 35.9 615 274 86 17 42 224.5 5.63 12.6 0.147 0.477 24.4 16.9 17.3 18.6 20.2 16.0 13.0 23.2 22.8 16.5 depress copper and 5 1CC1 - 1CC5 37.7 599 276 85 17 41 223.4 5.62 12.4 0.146 0.477 25.0 17.8 18.0 19.4 20.9 16.7 13.6 24.0 23.7 17.3 pyrite. Condition with 5 SVC 2.1 1363 45 129 28 75 354.0 6.99 26.7 0.161 0.452 3.1 0.2 1.5 1.7 2.1 1.5 0.9 2.8 1.4 0.9 nitrogen. 1CC + SVC 39.8 639 264 87 18 43 230.2 5.69 13.2 0.147 0.475 28.1 18.0 19.5 21.1 23.0 18.2 14.5 26.8 25.1 18.2 16 1CT (Average) 60.2 1088 788 238 44 95 681 22.1 23.7 0.288 1.42 71.9 82.0 80.5 78.9 77.0 81.8 85.5 73.2 74.9 81.8 1.04 Calc Head 909 579 178 34 74 501 15.6 19.5 0.232 1.04 8-2 n/a 59 1 10.0 1CC1 12.0 108 99 15 7 13 32 0.70 4.29 0.055 0.133 1.5 2.1 1.0 2.6 2.3 0.8 0.5 2.4 2.5 2.4 2 1CC1 - 1CC2 20.4 108 103 16 7 13 35 0.77 4.50 0.054 0.137 2.5 3.7 1.8 4.2 3.8 1.5 0.9 4.3 4.2 4.3 1. Condition with 3 1CC1 - 1CC3 24.6 120 116 19 7 14 42 0.92 5.08 0.058 0.148 3.4 5.0 2.6 5.7 5.1 2.2 1.3 5.8 5.4 3.0 activated DG-13 5 1CC1 - 1CC4 28.4 137 137 24 9 16 56 1.21 5.78 0.066 0.171 4.5 6.9 3.8 7.8 6.8 3.3 2.0 7.6 7.1 4.0 carbon followed by 5 1CC1 - 1CC5 29.9 154 151 28 10 18 66 1.40 6.55 0.071 0.184 5.3 8.0 4.7 9.2 8.1 4.1 2.4 9.1 8.1 4.5 depression of sulfides Apr-12 16 with sodium sulfide. 1CT (Average) 70.1 1193 737 248 42 90 674 24.33 28.3 0.347 1.68 94.7 92.0 95.3 90.8 91.9 95.9 97.6 90.9 91.9 95.5 1.06 Calc Head 882 561 182 32 68 492 17.5 21.8 0.264 1.23 8-3 n/a 59 1 10.0 1CC1 8.6 101 93 6 6 12 22 0.68 3.90 0.048 0.118 0.9 1.4 0.3 1.6 1.4 0.4 0.3 1.5 1.8 1.5 2 1CC1 - 1CC2 18.5 98 92 8 6 11 24 0.68 4.01 0.045 0.121 1.9 3.0 0.9 3.4 2.8 0.9 0.7 3.3 3.7 3.3 1. Repeat of Z10-8-2, 3 1CC1 - 1CC3 23.6 108 106 12 6 12 33 0.88 4.49 0.050 0.136 2.7 4.4 1.6 4.7 4.1 1.6 1.2 4.8 5.2 2.7 but addition of Cytec 5 1CC1 - 1CC4 26.7 130 125 17 8 15 47 1.11 5.34 0.058 0.153 3.7 5.9 2.6 6.5 5.6 2.6 1.7 6.5 6.8 3.5 404 to float bismuth 5 1CC1 - 1CC5 29.2 157 150 24 9 18 63 1.45 6.41 0.067 0.179 4.9 7.7 4.1 8.4 7.4 3.9 2.4 8.5 8.6 4.5 minerals. Target the Apr-12 16 Ag with the 404. 1CT (Average) 70.8 1270 747 235 42 94 643 24.5 28.4 0.296 1.58 95.1 92.3 95.9 91.6 92.6 96.1 97.6 91.5 91.4 95.5 1.06 Calc Head 944 572 173 32 71 473 17.8 22.0 0.229 1.17

Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Baseline Z10-6-1 1.23 1.22 1.04 1.13 1.13 Z10-8-1 1.19 1.04 1.06 1.08 1.11 Z10-8-2 1.03 1.06 1.02 1.07 1.06 Z10-8-3 1.03 1.06 1.02 1.07 1.05

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5.4. Depression of Bismuth Minerals – Final Concentrate

Since depression of the copper minerals did not provide a significant separation of bismuth minerals, testwork around the flotation of copper minerals while attempting to depress the bismuth minerals was initiated. As shown from the mineralogy of the final concentrate, liberation of both the copper and bismuth minerals was not a concern. Several parameters were tested and are summarized in Table 5.5.

Table 5.5: Summary of Metallurgical Bismuth Depression Screening Tests for Final Concentrate

Test ID Comments Z10- Baseline Test Series 8-13 Repeat of Z10-6-1 baseline test. Pre-Aeration Reagent Series 8-4 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and 7262 for bismuth depression. 8-5 Repeat of Z10-8-4, but with SO2 in place of air in preaeration. 8-6 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and 7261A for bismuth depression. 8-7 Repeat of Z10-8-6, but with SO2 in place of air in preaeration. Pre-Aeration Series 8-16 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and Nalco TX-15281 for bismuth depression. 10-1 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and Nalco TX-15155 for bismuth depression.

Several parameters were tested including the use of Cytec 7262 and 7261A to depress bismuth minerals at pH > 11.0, pre-aeration prior to flotation, and Nalco TX-15281 and TX-15155 for bismuth depression during flotation at pH > 11.0. Test Z10-8-13 was a repeat of the baseline that was completed during the testing of the bismuth depression parameters. Table 5.6 summarizes metallurgical results for the six bismuth depression methods trialed on the final concentrate. The bismuth reagent depression series was completed in conjunction with Baseline Test Z10-8-13 and baseline results are summarized in Table 5.6. A rejection ratio was calculated for each test completed in this series, based on the overall copper recovery divided by the overall bismuth recovery as used to compare results explained in previous sections. Baseline Test Z10-8-13 had a rejection ratio of 1.34; baseline Test Z10-6-1 was not included in Table 5.6 due to the time frame in which it was completed compared to the tests in this series. All tests in this series with the exception of the tests using the Nalco reagents, Z10-8-16 and Z10-10-1, had a better rejection ratio compared to the baseline. Test Z10-8-5 had the highest rejection ratio compared to all other tests in this series, 2.82, suggesting the parameters utilized in this test offer the best potential separation of the copper minerals from the bismuth minerals. Rejection ratios for each penalty element are provided at the bottom of Table 5.6. Compared to the baseline Test, Z10-8-5 had higher rejection ratios for all penalty elements with the exception of cadmium (Cd). The Baseline Test had a rejection ratio of 1.11 for Cd while Test Z10-8-5 was 1.09, which may be associated with analytical error or different mineral associations compared to other penalty elements. Cadmium was shown to almost entirely be associated with sphalerite where other penalty elements were associated between several different minerals which could explain why cadmium behaves differently than other penalty elements.

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Table 5.6: Metallurgical Results for Final Concentrate - Bismuth Mineral Depression Series

Test Test Ret. Assay Distribution, % Reject Test % Time Ratio Conditions Z10 Date Solids Product wt. As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn (mins) % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Final Concentrate Assayed Feed 100 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 8-13 5 8.0 1CC1 32.6 183 233 56 7 21 194 7.5 9.7 0.107 0.374 6.3 11.5 12.0 9.1 11.5 12.5 14.0 14.2 14.6 14.2 5 1CC1 - 1CC2 47.5 264 340 91 14 34 299 13.1 14.6 0.147 0.614 13.2 24.5 28.6 25.7 27.0 28.1 35.4 30.9 29.3 30.9 BASELINE 5 1CC1 - 1CC3 54.9 319 402 111 17 38 344 14.9 16.2 0.165 0.746 18.4 33.5 40.1 37.2 35.0 37.4 46.5 39.8 38.1 39.2 TEST/ FINAL 5 1CC1 - 1CC4 61.8 398 459 127 20 41 380 16.4 17.8 0.182 0.848 25.8 43.0 51.9 48.3 42.6 46.5 57.5 49.0 47.1 50.1 1.34

CONC 20 10-Oct-12 1CT (Average) 38.2 1856 980 191 34 90 709 19.51 29.9 0.330 1.37 74.2 57.0 48.1 51.7 57.4 53.5 42.5 51.0 52.9 49.9 Calc Head 955 658 152 25 60 506 17.6 22.4 0.239 1.05 8-4 5 10.0 1CC1 35.8 232 328 147 37 27 286 14.2 16.2 0.118 0.972 8.3 18.8 32.2 25.6 16.3 20.9 32.0 26.8 17.5 26.8 1. Pre-aeration, 5 1CC1 - 1CC2 40.7 241 331 143 36 27 282 13.8 16.0 0.120 0.948 9.8 21.6 35.6 28.6 18.2 23.4 35.5 30.2 20.3 30.2 11.0 w ith Cytec 5 1CC1 - 1CC3 48.0 293 354 152 37 30 309 14.7 17.2 0.130 0.989 14.0 27.2 44.6 34.0 24.1 30.3 44.7 38.3 26.1 44.2 7262/Cytec 3894. 5 1CC1 - 1CC4 50.4 324 366 153 37 31 316 14.8 17.4 0.136 0.996 16.3 29.5 47.3 35.9 26.3 32.4 47.2 40.7 28.5 46.7 1.60

Maintain pH 11.0 Oct-12 20 for flotation. 1CT (Average) 49.6 1684 886 174 67 88 666 16.8 25.8 0.346 1.15 83.7 70.5 52.7 64.1 73.7 67.6 52.8 59.3 71.5 53.3 Calc Head 998 624 163 52 59 489 15.8 21.6 0.240 1.072 8-5 5 10.0 1CC1 31.3 377 197 159 31 31 316 17.2 21.1 0.079 0.910 12.1 10.0 30.3 20.1 16.8 20.0 34.8 30.4 10.3 30.4 5 1CC1 - 1CC2 33.8 385 202 155 30 30 306 16.4 20.2 0.082 0.895 13.4 11.1 31.9 21.1 17.4 20.9 35.8 31.6 11.5 31.6 1. Repeat of Z10-8- 5 1CC1 - 1CC3 36.0 397 209 151 30 29 299 15.7 19.5 0.085 0.878 14.7 12.2 33.1 22.6 17.8 21.7 36.5 32.5 12.8 30.0 4, but w ith SO in 2 5 1CC1 - 1CC4 37.8 409 214 148 30 28 293 15.1 19.0 0.087 0.862 15.9 13.1 33.9 23.7 18.1 22.3 37.0 33.2 13.7 30.9 2.82 place of air for pre- aeration. Oct-12 20 1CT (Average) 62.2 1313 864 174 59 76 616 15.7 23.3 0.333 1.17 84.1 86.9 66.1 76.3 81.9 77.7 63.0 66.8 86.3 69.1 Calc Head 971 618 164 48 58 494 15.5 21.7 0.240 1.05 8-6 5 10.0 1CC1 38.0 275 308 155 22 32 342 15.1 18.3 0.145 0.918 10.6 19.3 35.6 17.1 20.5 26.3 33.8 31.6 23.5 31.6 5 1CC1 - 1CC2 41.8 273 314 152 23 31 336 14.6 17.7 0.147 0.905 11.6 21.7 38.5 19.8 22.1 28.4 35.9 33.8 26.3 33.8 1. Trial of Cytec 5 1CC1 - 1CC3 45.6 287 329 158 25 33 348 14.8 17.9 0.154 0.938 13.3 24.7 43.5 23.5 25.2 32.0 39.7 37.2 30.1 41.8 7261A w ith pre- 5 1CC1 - 1CC4 49.0 303 347 162 26 34 359 14.9 18.1 0.162 0.963 15.1 28.0 48.0 26.5 28.0 35.6 43.0 40.4 34.0 46.2 1.54

aeration. Oct-12 20 1CT (Average) 51.0 1648 858 169 71 84 623 19.0 25.7 0.303 1.08 84.9 72.0 52.0 73.5 72.0 64.4 57.0 59.6 66.0 53.8 Calc Head 988 607 166 49 59 493 17.0 22.0 0.234 1.02 8-7 5 10.0 1CC1 8.4 232 240 54 27 11 151 6.2 9.9 0.123 0.366 2.0 3.4 2.9 4.1 1.7 2.6 2.8 3.8 5.1 3.8 5 1CC1 - 1CC2 16.0 283 231 91 26 21 235 12.2 14.6 0.105 0.574 4.7 6.3 9.4 7.6 6.3 7.7 10.6 10.6 8.3 10.6 1. Repeat of test 5 1CC1 - 1CC3 19.8 290 231 95 25 20 230 11.9 14.2 0.104 0.587 6.0 7.8 12.1 9.0 7.3 9.3 12.8 12.8 10.1 11.3 Z10-8-6, but w ith 5 1CC1 - 1CC4 22.3 288 233 94 24 19 224 11.4 13.7 0.104 0.582 6.7 8.9 13.5 9.9 7.8 10.2 13.8 13.9 11.4 12.6 1.55

SO2. Oct-12 20 1CT (Average) 77.7 1153 691 172 64 65 568 20.5 24.5 0.232 1.16 93.3 91.1 86.5 90.1 92.2 89.8 86.2 86.1 88.6 87.4 Calc Head 959 589 154 55 54 491 18.5 22.1 0.203 1.03

Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Baseline Z10-8-13 2.23 1.34 1.11 1.19 1.35 Z10-8-4 2.90 1.60 1.00 1.31 1.79 Z10-8-5 2.33 2.82 1.09 1.56 2.04 Z10-8-6 2.85 1.54 0.90 1.62 1.54 Z10-8-7 2.06 1.55 1.02 1.39 1.77 Z10-8-16 1.04 1.05 1.01 1.01 1.04 Z10-10-1 1.04 1.05 1.00 1.02 1.03

53

After evaluating the rejection ratios of each test, Z10-8-5 resulted in the best separation between copper and bismuth minerals in this series of testwork. Flotation conditions for Test Z10-8-5 consisted of pre-aeration of the final concentrate for 30 minutes with sulfur dioxide (SO2) prior to flotation. After the 30 minutes had elapsed the pH of the slurry was adjusted to between pH 11.0 and 12.0 then conditioned with Cytec 3894 for the collection of copper minerals and Cytec 7262 for the depression of bismuth minerals. Test Z10-8-4 was similar to Test Z10-8-5, but was pre-aerated with air in place of SO2, which resulted in a lower rejection ratio of bismuth and other penalty elements. With the exception of the tests using Nalco reagents all separations were better than the baseline test for this series, Z10-8-13.

Figure 5.3 illustrates the cumulative copper grade recovery curve for this series of tests on the final concentrate. Based on Figure 5.3 it becomes apparent that the Nalco reagents used in Tests Z10-8- 16 and Z10-10-1 produced conditions that were beneficial for copper recovery and grade. Both Nalco reagents yield concentrates of higher copper grade and recovery compared to the baseline test and would be of interest for a further study as a replacement for the current copper collector utilized for these copper ores, but is beyond the scope of this study. Similar to the grade recovery curves shown in the final concentrate baseline series the copper grade increases with copper recovery. This is most likely due to the fact the testwork parameters operate in a depressant environment, causing an increase in copper mineral flotation over time. The coarse poorly liberated copper minerals float upfront or entrained in the froth with gangue and the finely fully liberated copper particles float slower. The MLA data shows that the fully liberated copper minerals occur at fine particle sizes, < 20 microns. Figure 5.4 shows the cumulative grade recovery curves for bismuth in the bismuth depression test series. Test Z10-8-5 shows the lowest bismuth grade and recovery compared to all other tests in this series, indicating that pre- aeration with SO2 and the use of 7262 lower the bismuth grade and recovery compared to the baseline test and the other tests in this series. Test Z10-8-7 is similar to Test Z10-8-5, but used Cytec 7261A in place of 7262 and yields slightly higher grade bismuth in the concentrate. Flotation tests using pre- aeration with Air rather than SO2 using both 7262 and 7261A offered no additional benefit as far as decreasing bismuth grade and recovery. It is also evident from Figure 5.4 that the Nalco reagents collect the bismuth minerals and achieved the highest bismuth grade and recovery in this test series. It becomes evident based on the graph presented in Figure 5.5 that the Nalco reagents didn’t facilitate a separation of copper from the bismuth. Both Nalco reagents seem to offer an increased ability to float the copper minerals while offering no separation from the bismuth minerals. Test Z10-8-5 shows a significant decrease in bismuth recovery for the same copper recovery when comparing the tests in this series indicating that a separation was made using the tests conditions. Both Tests Z10-8-4 and Z10-8-6 showed some separation when compared to the Baseline Test, but it was decided that the separation was not significant enough to explore further. Test Z10-8-7 yielded the lowest bismuth recovery, but also the lowest copper recovery suggesting over depressed flotation parameters. The conditions utilized in Test Z10-8-7 were not tested further.

54

Cumulative Copper Grade vs. Recovery 25.0

20.0

15.0

10.0 Cumulative Copper Grade, %

5.0

0.0 0.0 10.0 20.0 30.0 40.0 50.0 60.0 70.0 80.0 90.0 100.0 Cumultaive Copper Recovery, % Baseline Test Z10-8-4 Pre-aeration, 7262 Z10-8-5 Pre-aeration (SO2), 7262 Z10-8-6 Pre-aeration, 7261A Z10-8-7 Pre-aeration (SO2), 7261A Z10-8-16 Nalco TX-15281 Z10-10-1 Nalco TX-15155

Figure 5.3: Cumulative copper grade recovery graphs for bismuth depression test series on final concentrate

55

Cumulative Bismuth Grade vs. Recovery 800

700

600

500

400

300

Cumulative Bismuth, Grade, ppm 200

100

0 0.0 10.0 20.0 30.0 40.0 50.0 60.0 70.0 80.0 90.0 100.0 Cumulative Bismuth Recovery, % Baseline Test Z10-8-4 Pre-aeration, 7262 Z10-8-5 Pre-aeration (SO2), 7262 Z10-8-6 Pre-aeration 7261A Z10-8-7 Pre-aeration (SO2), 7261A Z10-8-16 Nalco TX-15281 Z10-10-1 Nalco Tx15155

Figure 5.4: Bismuth grade recovery graphs for bismuth depression test series on final concentrate

56

Cumulative Copper Recovery vs. Cumulative Bismuth

100.0 Recovery

90.0

80.0

70.0

60.0

50.0

40.0

30.0 Cumulative Bismuth Recovery, % 20.0

10.0

0.0 0.0 10.0 20.0 30.0 40.0 50.0 60.0 70.0 80.0 90.0 100.0 Cumulative Copper Recovery % Baseline Test Z10-8-4 Pre-aeration, 7262 Z10-8-5 Pre-aeration (SO2), 7262 Z10-8-6 Pre-aeration, 7261A Z10-8-7 Pre-aeration (SO2), 7261A Z10-8-16 Nalco TX-15281 Z10-10-1 Nalco TX-15155

Figure 5.5 Cumulative copper recovery vs. cumulative bismuth recovery for bismuth depression series on final concentrate

57

Cumulative Copper Grade vs. Cumulative Bismuth Grade

800

700

600

500

400

300

Cumulative Bismuth Grade, ppm 200

100

0 0.0 5.0 10.0 15.0 20.0 25.0 Cumulative Copper Grade % Baseline Test Z10-8-4 Pre-aeration, 7262 Z10-8-5 Pre-aeration (SO2), 7262 Z10-8-6 Pre-aeration, 7261A Z10-8-7 Pre-aeration (SO2), 7261A Z10-8-16 Nalco TX-15281 Z10-10-1 Nalco TX-15155

Figure 5.6 Cumulative copper grade vs. cumulative bismuth grade for bismuth depression series on final concentrate

Figure 5.6 shows cumulative copper grade as a function of cumulative bismuth grade for the tests in this series. The Nalco reagents produced copper grades that were higher than the baseline, but bismuth grades were also higher compared to the Baseline Test. The Nalco reagents behave differently to the other reagents used in the test series. The purpose was to create a depressant environment to separate copper from the bismuth, where the Nalco reagents were shown activate both the copper and bismuth minerals offering no separation. When comparing cumulative copper grade to cumulative bismuth grade, Test Z10-8-5 resulted in the lowest bismuth grade while still obtaining reasonable copper grade, although improvement of copper recovery would be a necessary objective for further testwork. Similar to the results shown in Figure 5.5, both Tests Z10-8-4 and Z10-8-6 resulted in slightly better results than the Baseline Test, but again the separation was not considered significant enough to explore further testing.

5.5. Baseline Metallurgical Testwork Results- Rougher Concentrate

Two baseline metallurgical tests were completed on the rougher concentrate that utilized the same baseline flotation parameters completed on the final concentrate. The flotation tests were completed using the same flotation retention times, but without the use of regrinding for surface preparation. Table 5.7 summarizes metallurgical results that were achieved from both baseline flotation tests completed on the rougher concentrate. The assayed feed for the rougher concentrate is provided in Table 5.7 and can

58 be compared to the calculated assays for each element. Both copper and lead assayed values are in good agreement with the calculated values giving confidence in the metallurgical balances for both tests. The majority of the other elements have relatively decent accountabilities.

Baseline metallurgical results for the rougher concentrate were similar to those achieved in baseline tests on the final concentrate. The Cu/Bi rejection ratios for the final concentrate were 0.88 and 1.34 and the rougher concentrate baseline tests resulted in rejection ratios of 1.18 and 1.29 respectively. In both baseline tests on the rougher concentrate the calculated heads for all elements were very similar, giving confidence in the baseline test results for comparison against reagent screening tests. However, the arsenic ratios between each baseline test were slightly different, suggesting the arsenic minerals are more sensitive to aging most likely the arsenopyrite (Bruckard 2007). Figures 5.7 and 5.8 illustrate the cumulative grade recovery curves for copper and bismuth on the rougher concentrate baseline tests. It is important to note the typical grade recovery curve relationship for the rougher concentrate baseline test series was observed which differed from the results explained for the final concentrate baseline test series.

Cumulative Copper Grade vs. Recovery

20.0

18.0

16.0

14.0

12.0

10.0

8.0

6.0 Cumulative Copper Grade,%

4.0

2.0

0.0 0.0 10.0 20.0 30.0 40.0 50.0 60.0 70.0 80.0 90.0 100.0 Cumulative Copper Recovery, % Baseline 3/9/2012 Baseline 12/20/2012

Figure 5.7: Cumulative copper grade recovery curve for rougher concentrate baseline tests

59

Table 5.7: Baseline Metallurgical Results for Rougher Concentrate Flotation Tests

Ret. Test Assay Distribution, % Reject Test Test Time % Regrind P80 Ratio Conditions Z10 Date mins. Solids Product wt. As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Rougher Concentrate Assayed Head 100 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 8-12 n/a 62 5 9.0 1CC1 43.1 13567 752 353 45 86 448 17.0 35.9 0.398 2.81 35.9 34.3 51.4 35.2 42.8 40.8 55.1 48.6 40.2 48.6 5 1CC1 - 1CC2 55.8 14808 834 356 48 91 459 16.2 36.2 0.429 2.89 50.8 49.3 67.2 48.7 58.7 54.2 68.1 63.7 56.2 63.7 Baseline 5 1CC1 - 1CC3 65.2 15736 895 352 52 92 473 15.8 36.8 0.441 2.98 63.1 61.8 77.6 61.5 69.3 65.3 77.5 75.6 67.6 63.5 Rougher 5 1CC1 - 1CC4 71.8 16278 932 349 54 92 479 15.4 36.7 0.443 3.05 71.9 70.9 84.6 70.8 76.5 72.8 83.5 82.9 74.8 71.6 1.18

Concentrate 20 9-Mar-12 1CT (Average) 28.2 16201 972 162 57 72 456 7.72 19.2 0.384 3.09 28.1 29.1 15.4 29.2 23.5 27.2 16.5 17.1 25.2 28.4 Calc Head 16256 943 296 55 87 473 13.3 31.7 0.426 3.06 10-2 n/a 62 5 9.0 1CC1 43.0 6308 705 389 38 120 389 18.9 37.0 0.329 3.25 16.4 32.7 59.3 27.5 42.1 36.1 60.1 50.3 33.5 50.3 5 1CC1 - 1CC2 54.5 7437 803 376 43 125 413 17.5 36.9 0.377 3.23 24.5 47.2 72.6 39.3 55.5 48.6 70.8 63.7 48.7 63.7 Baseline 5 1CC1 - 1CC3 61.6 8345 862 365 47 129 428 16.9 36.8 0.399 3.24 31.1 57.3 79.7 48.4 65.0 56.9 77.0 71.7 58.2 62.5 Rougher 5 1CC1 - 1CC4 65.2 8821 886 360 48 131 434 16.6 36.7 0.407 3.26 34.8 62.3 83.2 52.9 69.4 61.1 80.1 75.7 62.7 66.6 1.29 Concentrate 20 20-Dec-12 1CT (Average) 34.8 30919 1007 137 81 108 518 7.74 22.1 0.450 3.07 65.2 37.7 16.8 47.1 30.6 38.9 19.9 24.3 37.3 33.4 Calc Head 16519 928 282 59 123 463 13.5 31.6 0.422 3.20

Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Baseline Z10-8-12 1.16 1.18 0.99 1.18 1.09 Baseline Z10-10-2 2.30 1.29 0.96 1.51 1.15

60

Cumulative Bismuth Grade vs. Recovery 1000

900

800

700

600

500

400

300

Cumulative Bismuth Grade,ppm 200

100

0 0.0 10.0 20.0 30.0 40.0 50.0 60.0 70.0 80.0 90.0 100.0 Cumulative Bismuth Recovery, % Baseline 3/9/2012 Baseline 12/20/2012

Figure 5.8: Cumulative bismuth grade recovery curves for rougher concentrate baseline tests

In Figures 5.7 and 5.8 the cumulative grade recovery curves for copper and bismuth are shown, and indicate that high cumulative grades and recoveries are achieved even in the absence of additional reagents. This suggests that both the copper and bismuth in the rougher concentrate are floatable with only the residual reagents from the process plant present in the sample or the copper and minerals present in the rougher concentrate are of coarser size. Bismuth cumulative grade and recovery increase with time, which could be due to the fine particle sizes of the bismuth minerals and their slow flotation kinetics over time. Based on the figures presented there can be a great deal of confidence in both baseline tests and the reproducibility of the results.

5.6. Depression of Bismuth Minerals – Rougher Concentrate

Some success in separation testwork on the final concentrate was achieved therefore the separation of copper from bismuth in the rougher concentrate was evaluated to validate the previous result and evaluate whether a similar or improved separation can be made on this material. The parameters utilized in the depression of the bismuth minerals in the final concentrate were utilized in the testing of the rougher concentrate. Testwork conditions included the trial of Cytec 7262 and 7261A as well as Nalco TX-15281 and TX-15155 reagents for penalty element depression. The combination of penalty element

61 depression reagents with pre-aeration using air and air/SO2 were also used to evaluate the response on the rougher concentrate compared to the final concentrate. Table 5.8 provides a brief summary of testwork completed on the rougher concentrate for the bismuth depression reagent series.

Table 5.8: Summary of Metallurgical Bismuth Depression Reagent Tests for Rougher Concentrate

Rougher Concentrate Baseline Series 8-12 Baseline flotation test, no reagents. Natural pH 10-2 Repeat of baseline flotation test, no reagents. Natural pH Pre-Aeration Reagent Series 8-8 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and Nalco TX-15155 for bismuth depression. 8-9 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and Nalco TX-15281 for bismuth depression. 8-10 Repeat of Z10-8-9, but with SO2 in place of air in preaeration. 8-11 Repeat of Z10-8-8, but with SO2 in place of air in preaeration. 8-14 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and 7261A for bismuth depression. 8-15 No regrind, preaeration. Increase pH to 11.0, Cytec 3894 for copper and 7262 for bismuth depression. 10-5 SO2 preaeration with no reagents. 10-6 Repeat of Z10-8-14, but with SO2 in place of air in preaeration.

Metallurgical results for the rougher concentrate are provided in Tables 5.9 and 5.10. The majority of the results were consistent with those shown in the final concentrate testwork. The superior test conditions for the final concentrate utilized pre-aeration with SO2 (30 minutes), adjustment to pH >11.0 with hydrated lime, Cytec 3894 for copper flotation and Cytec 7262 for bismuth depression. Testwork completed on the rougher concentrate resulted in a rejection ratio of 3.66 for Test Z10-10-4, which was the best for this series of tests. Below Table 5.9 are the calculated rejection ratios for each of the penalty elements, again Test Z10-10-4 had better rejections ratios for all penalty elements compared to results from baseline tests.

Rejection ratios for arsenic and antimony for Test Z10-10-4 were ~2.5 times better than those in the baseline tests. The trend of penalty element depression occurring for all elements suggests the penalty elements studied in this test program behave similarly during flotation. There are some exceptions to this trend, cadmium and arsenic. Cadmium has been shown to be virtually entirely associated with sphalerite, whereas other penalty elements could be found in several different minerals suggesting that cadmium that is associated with sphalerite might require a different separation approach. The arsenic also separates better in the rougher concentrate compared to the final concentrate. This could be due to a larger majority of the arsenic being associated with arsenopyrite in the rougher concentrate than in the final concentrate. Enargite content was also shown to be higher in content in the final concentrate compared to the rougher concentrate. It has been well documented that arsenopyrite can be depressed at elevated pH and with pre-aeration, but enargite shows more resistance to pH (Ma 2009).

62

Table 5.9: Metallurgical Results for Rougher Concentrate Flotation Tests

Ret. Test Assay Distribution, % Reject Regr Test Time % Solids Test Conditions P80 Ratio Z10 ind Date mins. Product wt. As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Rougher Concentrate Assayed Head 100 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 8-12 n/a 62 5 9.0 1CC1 43.1 13567 752 353 45 86 448 17.0 35.9 0.398 2.81 35.9 34.3 51.4 35.2 42.8 40.8 55.1 48.6 40.2 48.6 5 1CC1 - 1CC2 55.8 14808 834 356 48 91 459 16.2 36.2 0.429 2.89 50.8 49.3 67.2 48.7 58.7 54.2 68.1 63.7 56.2 63.7 Baseline 5 1CC1 - 1CC3 65.2 15736 895 352 52 92 473 15.8 36.8 0.441 2.98 63.1 61.8 77.6 61.5 69.3 65.3 77.5 75.6 67.6 63.5 Rougher 5 1CC1 - 1CC4 71.8 16278 932 349 54 92 479 15.4 36.7 0.443 3.05 71.9 70.9 84.6 70.8 76.5 72.8 83.5 82.9 74.8 71.6 1.18

Concentrate 20 9-Mar-12 1CT (Average) 28.2 16201 972 162 57 72 456 7.72 19.2 0.384 3.09 28.1 29.1 15.4 29.2 23.5 27.2 16.5 17.1 25.2 28.4 Calc Head 16256 943 296 55 87 473 13.3 31.7 0.426 3.06 10-2 n/a 62 5 9.0 1CC1 43.0 6308 705 389 38 120 389 18.9 37.0 0.329 3.25 16.4 32.7 59.3 27.5 42.1 36.1 60.1 50.3 33.5 50.3 5 1CC1 - 1CC2 54.5 7437 803 376 43 125 413 17.5 36.9 0.377 3.23 24.5 47.2 72.6 39.3 55.5 48.6 70.8 63.7 48.7 63.7 Baseline 5 1CC1 - 1CC3 61.6 8345 862 365 47 129 428 16.9 36.8 0.399 3.24 31.1 57.3 79.7 48.4 65.0 56.9 77.0 71.7 58.2 62.5 Rougher 5 1CC1 - 1CC4 65.2 8821 886 360 48 131 434 16.6 36.7 0.407 3.26 34.8 62.3 83.2 52.9 69.4 61.1 80.1 75.7 62.7 66.6 1.29 Concentrate 20 20-Dec-12 1CT (Average) 34.8 30919 1007 137 81 108 518 7.74 22.1 0.450 3.07 65.2 37.7 16.8 47.1 30.6 38.9 19.9 24.3 37.3 33.4 Calc Head 16519 928 282 59 123 463 13.5 31.6 0.422 3.20 1. Trial of Nalco 8-8 n/a 62 5 11.0 1CC1 35.1 4985 1175 602 82 96 703 27.7 32.4 0.620 4.38 10.7 47.3 68.6 31.3 35.5 52.3 70.3 35.5 56.5 35.5 TX 15155 on 5 1CC1 - 1CC2 45.9 7861 1190 540 97 106 707 25.2 32.9 0.588 4.32 22.0 62.5 80.2 48.5 51.1 68.7 83.5 47.1 69.9 47.1 rougher 5 1CC1 - 1CC3 51.4 9575 1183 508 101 109 703 24.1 33.2 0.568 4.30 30.1 69.7 84.7 56.6 59.0 76.6 89.6 53.4 75.7 78.0 concentrate. 5 1CC1 - 1CC4 53.8 10499 1181 496 103 110 701 23.5 33.3 0.561 4.29 34.5 72.7 86.4 60.2 62.4 79.8 91.6 55.9 78.2 81.3 1.26

Pre-aeration prior Mar-12 20 to reagent 1CT (Average) 46.2 23273 515 91 80 78 207 2.53 30.6 0.183 1.15 65.5 27.3 13.6 39.8 37.6 20.2 8.4 44.1 21.8 18.7 additions. Calc Head 16405 873 308 92 95 473 13.8 32.0 0.386 2.84 1. Trial of Nalco 8-9 n/a 62 5 11.0 1CC1 36.3 5614 1141 598 107 135 707 28.2 32.4 0.605 4.36 12.3 47.4 69.7 39.6 45.6 53.8 72.4 60.4 56.4 60.4 TX 15281on 5 1CC1 - 1CC2 46.0 8204 1158 542 112 134 705 26.0 33.1 0.575 4.28 22.8 60.9 80.1 52.4 57.3 67.9 84.6 78.2 67.8 78.2 rougher 5 1CC1 - 1CC3 51.6 9740 1146 509 114 133 695 24.8 33.4 0.551 4.22 30.4 67.6 84.3 60.0 63.8 75.0 90.2 88.3 72.9 76.5 concentrate. 5 1CC1 - 1CC4 54.6 10932 1150 495 116 134 693 24.0 33.4 0.545 4.22 36.1 71.8 86.6 64.4 67.8 79.2 92.5 93.6 76.4 81.1 1.29

Pre-aeration prior Mar-12 20 to reagent 1CT (Average) 45.4 23301 543 92 77 77 220 2.33 2.77 0.202 1.19 63.9 28.2 13.4 35.6 32.2 20.8 7.5 6.4 23.6 18.9 additions. Calc Head 16543 875 312 98 108 478 14.2 19.5 0.390 2.84 1. Trial of Nalco 8-10 n/a 62 5 11.0 1CC1 20.9 4199 632 487 12 54 690 27.9 28.7 0.493 2.97 5.2 13.5 29.0 4.2 15.3 29.1 44.8 19.0 22.9 19.0 TX 15281on 5 1CC1 - 1CC2 27.6 4876 767 582 22 60 710 26.8 29.1 0.546 3.64 8.0 21.7 45.7 10.2 22.5 39.6 56.9 25.5 33.6 25.5 rougher 5 1CC1 - 1CC3 33.6 5891 1010 683 39 74 740 25.4 29.4 0.637 4.25 11.7 34.8 65.3 22.3 33.9 50.3 65.7 31.3 47.7 46.9 concentrate. 5 1CC1 - 1CC4 36.4 6538 1086 684 47 79 752 24.6 29.5 0.659 4.35 14.1 40.6 70.9 28.6 38.9 55.4 69.0 34.1 53.6 52.0 1.70

Pre-aeration with Dec-12 20

SO2 prior to 1CT (Average) 63.6 22960 912 161 67 71 347 6.32 32.6 0.327 2.30 85.9 59.4 29.1 71.4 61.1 44.6 31.0 65.9 46.4 48.0 reagent Calc Head 16983 975 351 59 74 494 13.0 31.5 0.448 3.04 Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Baseline 8-12 1.16 1.18 0.99 1.18 1.09 Baseline 10-2 2.30 1.29 0.96 1.51 1.15 Z10-8-8 2.66 1.26 1.06 1.52 1.47 Z10-8-9 2.56 1.29 1.07 1.44 1.36 Z10-8-10 4.89 1.70 0.97 2.41 1.77

63

Table 5.10: Metallurgical Results for Rougher Concentrate Flotation Tests

Ret. Test Assay Distribution, % Reject Regr Test Time % Solids Test Conditions P80 Ratio Z10 ind Date mins. Product wt. As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Rougher Concentrate Assayed Head 100 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 1. Trial of Nalco 8-11 n/a 62 5 11.0 1CC1 13.8 2492 586 757 22 60 544 28.7 29.5 0.371 5.41 2.0 8.0 31.0 5.0 10.4 15.3 28.9 12.6 11.7 12.6 TX 15155 on 5 1CC1 - 1CC2 18.2 3061 731 726 30 69 576 28.0 29.3 0.438 5.19 3.3 13.2 39.3 8.9 15.7 21.4 37.3 16.5 18.3 16.5 rougher 5 1CC1 - 1CC3 29.0 3972 979 740 43 84 671 27.1 29.6 0.561 5.28 6.8 28.1 63.5 20.5 30.6 39.6 57.4 26.4 37.2 48.1 concentrate. 5 1CC1 - 1CC4 38.5 6053 1159 662 57 97 724 25.8 30.4 0.651 5.03 13.7 44.1 75.4 35.8 46.7 56.8 72.6 36.1 57.2 60.8 1.65

Pre-aeration with Mar-12 20

SO2 prior to 1CT (Average) 61.6 23768 919 135 64 69 343 6.08 33.6 0.306 2.02 86.3 55.9 24.6 64.2 53.3 43.2 27.4 63.9 42.8 39.2 reagent Calc Head 16956 1011 338 61 80 489 13.7 32.4 0.438 3.18 8-14 n/a 62 5 9.0 1CC1 20.3 4362 915 709 44 74 624 26.3 31.3 0.557 5.37 5.4 19.2 48.0 14.2 16.3 25.5 41.2 19.7 26.1 19.7 5 1CC1 - 1CC2 31.7 5798 1059 625 54 87 699 25.8 30.8 0.618 5.05 11.2 34.8 66.2 27.3 29.8 44.7 63.3 30.3 45.3 30.3 Trial of Cytec 5 1CC1 - 1CC3 33.7 6310 1093 610 57 88 706 25.1 30.6 0.629 5.03 12.9 38.1 68.6 30.3 32.2 47.9 65.5 31.9 48.9 57.8 7261A, pH 11.0. 5 1CC1 - 1CC4 35.1 6750 1114 598 58 89 708 24.6 30.5 0.632 5.01 14.4 40.5 70.2 32.3 33.9 50.1 66.9 33.1 51.2 60.0 1.65 Rougher Dec-12 20 Concentrate 1CT (Average) 64.9 21729 882 138 66 94 381 6.62 33.2 0.326 1.80 85.6 59.5 29.8 67.7 66.1 49.9 33.1 66.9 48.8 40.0 Calc Head 16471 963 299 63 92 495 12.9 32.2 0.433 2.93 8-15 n/a 62 5 9.0 1CC1 15.7 3392 822 795 27 85 587 26.6 28.9 0.509 6.19 3.2 13.4 41.1 7.2 14.5 18.7 31.1 14.1 18.3 14.1 5 1CC1 - 1CC2 24.0 4322 926 696 34 87 665 26.9 29.2 0.560 5.57 6.2 23.1 55.0 13.7 22.8 32.4 48.0 21.8 30.8 21.8 Trial of Cytec 5 1CC1 - 1CC3 28.6 4987 969 658 38 90 689 26.6 29.3 0.583 5.39 8.6 28.9 62.1 18.7 28.0 40.1 56.7 26.1 38.3 51.8 7262, pH 11.0. 5 1CC1 - 1CC4 32.6 5750 1019 623 43 93 696 26.0 29.5 0.601 5.24 11.3 34.6 67.0 23.7 33.1 46.2 63.1 29.9 44.9 57.3 1.82 Rougher Dec-12 20 Concentrate 1CT (Average) 67.4 21739 931 149 66 91 393 7.35 33.4 0.356 1.88 88.7 65.4 33.0 76.3 66.9 53.8 36.9 70.1 55.1 42.7 Calc Head 16532 960 303 58 92 491 13.4 32.2 0.436 2.98 10-4 n/a 62 5 9.0 1CC1 14.1 3173 331 173 10 85 395 28.2 27.8 0.142 1.84 2.8 4.9 10.6 2.3 11.2 12.7 30.2 13.4 4.7 13.4 5 1CC1 - 1CC2 24.1 3763 413 251 16 87 416 26.9 27.5 0.162 2.52 5.7 10.5 26.5 6.3 19.7 22.9 49.5 22.8 9.2 22.8 Pre-aeration with 5 1CC1 - 1CC3 26.3 4348 463 258 19 88 425 26.1 27.3 0.188 2.66 7.2 12.8 29.7 8.4 21.8 25.5 52.4 24.7 11.6 23.7 SO2 and Cytec 5 1CC1 - 1CC4 27.8 4867 504 257 22 89 432 25.4 27.1 0.207 2.72 8.5 14.7 31.3 10.1 23.3 27.4 53.8 25.9 13.5 25.7 3.66 7262. 3894 for Dec-12 20 Cu. 1CT (Average) 72.2 20138 1129 218 75 114 440 8.39 29.9 0.514 3.03 91.5 85.3 68.7 89.9 76.7 72.6 46.2 74.1 86.5 74.3 Calc Head 15892 955 229 60 107 438 13.1 29.1 0.428 2.95 10-6 n/a 62 5 9.0 1CC1 4.3 12929 1161 242 64 100 502 12.5 22.2 0.554 3.59 3.5 5.4 4.7 4.3 4.1 5.1 4.0 3.2 5.3 3.2 5 1CC1 - 1CC2 6.8 12520 1130 246 60 100 507 13.8 22.9 0.537 3.59 5.3 8.3 7.5 6.4 6.4 8.1 7.0 5.2 8.1 5.2 5 1CC1 - 1CC3 9.1 12802 1150 250 62 101 515 13.9 23.2 0.544 3.66 7.2 11.2 10.1 8.7 8.6 10.9 9.4 7.0 10.9 11.1 Preaeration with 5 1CC1 - 1CC4 10.9 13159 1167 251 62 102 519 13.9 23.6 0.552 3.73 8.9 13.6 12.1 10.5 10.4 13.2 11.3 8.5 13.3 13.6 0.83 SO2 and 7261A. Dec-12 20 1CT (Average) 89.1 16598 895 222 64 107 420 13.39 31.0 0.442 2.88 91.1 86.4 87.9 89.5 89.6 86.8 88.7 91.5 86.7 86.4 Calc Head 16223 925 225 64 106 431 13.4 30.2 0.454 2.97

Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Z10-8-11 5.30 1.65 0.96 2.03 1.55 Z10-8-14 4.65 1.65 0.95 2.07 1.97 Z10-8-15 5.58 1.82 0.94 2.66 1.91 Z10-10-4 6.33 3.66 1.72 5.33 2.31 Z10-10-6 1.27 0.83 0.93 1.08 1.09

64

Figure 5.9 shows the cumulative copper grade recovery curves for this bismuth depression test series completed on the rougher concentrate. All tests with the exception of Z10-10-6 have improved the grade recovery curves compared to the two baseline tests. Again, the Nalco reagents suggest that they have great affinity for copper minerals due to the improved grade recovery curves compared to the other tests in this series.

Cumulative Copper Grade vs. Recovery 35.0

30.0

25.0

20.0

15.0

10.0 Cumulative Copper Grade,% 5.0

0.0 0.0 20.0 40.0 60.0 80.0 100.0 Cumulative Copper Recovery, % Z10-8-8 Pre-aeration, TX-15155 Z10-8-9 Pre-aeration, TX-15281 Z10-8-11 re-aeration (SO2), TX-15155 Z10-8-10 Pre-aeration (SO2), TX-15281 Z10-8-14 7261A Z10-8-15 7262 Z10-10-4 Pre-aeration (SO2), 7262 Z10-10-6 Pre-aeration (SO2), 7261A

Figure 5.9: Cumulative copper grade recovery curves for rougher concentrate bismuth depression series

Figure 5.10 shows the cumulative grade recovery curves for bismuth for all tests completed in this series of testwork on the rougher concentrate. Again, the baseline tests are very similar suggesting great confidence in the results obtained in the test series. The only exception would be arsenic, which suggests the arsenic minerals are more sensitive to aging as discussed in Section 5.5. All reagent conditions in this series with the exception of Test Z10-10-4 improved the grade recovery curves for bismuth compared to baseline tests, which was not the goal of testwork program. The conditions utilized in Test Z10-10-4 suggest that the set of conditions can still improve the copper performance (from a cumulative copper grade perspective) shown Figure 5.10, while limiting the bismuth flotation performance. Additional work will need to be pursued in order to improve overall copper recoveries under

65 these conditions. Since this test series targeted the depression of bismuth minerals the grade recovery curve shown for the test series was not of concern, but confirms that additional bismuth will float with time. The supplementary bismuth floating suggests additional reagent might be needed to depress the bismuth.

Cumulative Bismuth Grade vs. Recovery 1400

1200

1000

800

600

400

Cumulative Bismuth Grade,ppm 200

0 0.0 20.0 40.0 60.0 80.0 100.0 Cumulative Bismuth Recovery, % Baseline Test Mar 2012 Baseline Test Dec 2012 Z10-8-8 Pre-aeration, TX-15155 Z10-8-9 Pre-aeration, TX-15281 Z10-8-11 re-aeration (SO2), TX-15155 Z10-8-10 Pre-aeration (SO2), TX-15281 Z10-8-14 7261A Z10-8-15 7262 Z10-10-4 Pre-aeration (SO2), 7262 Z10-10-6 Pre-aeration (SO2), 7261A

Figure 5.10: Cumulative bismuth grade recovery curves for bismuth depression test series on rougher concentrate

Figure 5.11 compares the cumulative copper grade to the cumulative bismuth grade for each test completed in the bismuth depression test series on the rougher concentrate. Since one of the goals of this testwork was to keep copper grade high while limiting the bismuth grade it was import to look at this comparison. Both baseline tests yielded similar results, while the majority of other tests didn’t seem to offer a decrease in bismuth grade although achieving higher copper grades than baseline tests. However, Test Z10-10-4 generated a final concentrate that was higher in copper grade and lower in bismuth grade compared to baseline tests. Test Z10-10-4 achieved ~25 % copper grade while maintaining the lowest bismuth grade, 504 ppm Bi. The results suggest that the conditions utilized in Z10-10-4 will separate the copper and bismuth minerals in the rougher concentrate, but may require

66 future optimization of flotation parameters to increase copper recovery. The pre-aeration with SO2, pH 11.0, and the use of Cytec 7262 offered the best separations of bismuth from copper in both the final and rougher concentrates.

Cumulative Copper Grade vs. Cumulative Bismuth Grade 1400

1200

1000

800

600

400

200 Cumulative Bismuth Grade,ppm

0 0.0 5.0 10.0 15.0 20.0 25.0 30.0 35.0 Cumulative Copper Grade, % Baseline Test Mar 2012 Baseline Test Dec 2012 Z10-8-8 Pre-aeration, TX-15155 Z10-8-9 Pre-aeration, TX-15281 Z10-8-11 re-aeration (SO2), TX-15155 Z10-8-10 Pre-aeration (SO2), TX-15281 Z10-8-14 7261A Z10-8-15 7262 Z10-10-4 Pre-aeration (SO2), 7262 Z10-10-6 Pre-aeration (SO2), 7261A

Figure 5.11 Cumulative copper grade vs. cumulative bismuth grade for bismuth depression test series on rougher concentrate

In Figure 5.12 when cumulative copper and bismuth recoveries are plotted against each other it becomes evident that the separation in Test Z10-10-4 was significant compared to all other tests in the series. All tests achieved higher copper recoveries compared to baseline tests, but also had high bismuth recoveries resulting in little to no separation between copper and bismuth. As stated previously, Test Z10-10-4 showed the most promising results for copper and bismuth separation and additional work is recommended to optimize copper recovery. In operation, there is the possibility of and/or regrinding of the flotation tails from this product, which may assist in further improvements in copper recovery. Addition of copper selective collectors may be another option that could be used to improve recoveries. With the identification of this bismuth depression chemistry, further testwork evaluating the impact of pH, pre-aeration, collector dosage etc. would be worthwhile exploring on a variety of copper ore types, utilizing fresh products from the copper flotation mill.

67

Cumulative Copper Recovery vs. Cumulative Bismuth Recovery 80.0

70.0

60.0

50.0

40.0

30.0

20.0

10.0 Cumulative Bismuth Reocvery, %

0.0 0.0 20.0 40.0 60.0 80.0 100.0 Cumulative Copper Recovery, % Baseline Test Mar 2012 Baseline Test Dec 2012 Z10-8-8 Pre-aeration, TX-15155 Z10-8-9 Pre-aeration, TX-15281 Z10-8-11 re-aeration (SO2), TX-15155 Z10-8-10 Pre-aeration (SO2), TX-15281 Z10-8-14 7261A Z10-8-15 7262 Z10-10-4 Pre-aeration (SO2), 7262 Z10-10-6 Pre-aeration (SO2), 7261A

Figure 5.12 Cumulative copper recoveries vs. cumulative bismuth recoveries for bismuth depression test series on rougher concentrate

68

Table 5.11: Metallurgical Results for Rougher Concentrate Pre-aeration Flotation Tests

Test Assay Distribution, % Reject Test Test Ret. % Z10 Regrind P80 Product wt As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn Ratio Conditions Date Time Solids % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Rougher Concentrate Assayed Head 100 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 8-12 n/a 62 5 9.0 1CC1 43.1 13567 752 353 45 86 448 17.0 35.9 0.398 2.81 35.9 34.3 51.4 35.2 42.8 40.8 55.1 48.6 40.2 48.6 5 1CC1 - 1CC2 55.8 14808 834 356 48 91 459 16.2 36.2 0.429 2.89 50.8 49.3 67.2 48.7 58.7 54.2 68.1 63.7 56.2 63.7 Baseline 5 1CC1 - 1CC3 65.2 15736 895 352 52 92 473 15.8 36.8 0.441 2.98 63.1 61.8 77.6 61.5 69.3 65.3 77.5 75.6 67.6 63.5 Rougher 5 1CC1 - 1CC4 71.8 16278 932 349 54 92 479 15.4 36.7 0.443 3.05 71.9 70.9 84.6 70.8 76.5 72.8 83.5 82.9 74.8 71.6 1.18

Concentrate 20 9-Mar-12 1CT (Average) 28.2 16201 972 162 57 72 456 7.72 19.2 0.384 3.09 28.1 29.1 15.4 29.2 23.5 27.2 16.5 17.1 25.2 28.4 Calc Head 16256 943 296 55 87 473 13.3 31.7 0.426 3.06 10-2 n/a 62 5 9.0 1CC1 43.0 6308 705 389 38 120 389 18.9 37.0 0.329 3.25 16.4 32.7 59.3 27.5 42.1 36.1 60.1 50.3 33.5 50.3 5 1CC1 - 1CC2 54.5 7437 803 376 43 125 413 17.5 36.9 0.377 3.23 24.5 47.2 72.6 39.3 55.5 48.6 70.8 63.7 48.7 63.7 Baseline 5 1CC1 - 1CC3 61.6 8345 862 365 47 129 428 16.9 36.8 0.399 3.24 31.1 57.3 79.7 48.4 65.0 56.9 77.0 71.7 58.2 62.5 Rougher 5 1CC1 - 1CC4 65.2 8821 886 360 48 131 434 16.6 36.7 0.407 3.26 34.8 62.3 83.2 52.9 69.4 61.1 80.1 75.7 62.7 66.6 1.29 Concentrate 20 20-Dec-12 1CT (Average) 34.8 30919 1007 137 81 108 518 7.74 22.1 0.450 3.07 65.2 37.7 16.8 47.1 30.6 38.9 19.9 24.3 37.3 33.4 Calc Head 16519 928 282 59 123 463 13.5 31.6 0.422 3.20 10-3 n/a 62 5 8.0 1CC1 37.0 5495 851 457 42 133 456 20.3 35.0 0.459 3.69 12.5 34.1 59.2 25.8 40.7 36.6 56.5 41.4 42.5 41.4 5 1CC1 - 1CC2 50.6 6867 911 416 47 132 465 18.5 35.8 0.459 3.35 21.4 50.0 73.8 39.4 55.5 51.1 70.4 57.9 58.2 57.9 Pre-aeration 5 1CC1 - 1CC3 57.8 7607 933 396 48 133 468 17.6 35.9 0.453 3.32 27.1 58.5 80.2 46.3 63.6 58.7 76.8 66.5 65.6 62.5 (Air) with no 5 1CC1 - 1CC4 63.6 8351 945 380 49 132 472 16.9 35.8 0.448 3.31 32.7 65.2 84.7 52.3 69.7 65.1 81.2 72.9 71.5 68.6 1.25

reagents Dec-12 20 1CT (Average) 36.4 30066 884 120 79 101 442 6.82 23.2 0.312 2.65 67.3 34.8 15.3 47.7 30.3 34.9 18.8 27.1 28.5 31.4 Calc Head 16256 923 285 60 121 461 13.2 31.2 0.399 3.07 10-5 n/a 62 5 8.0 1CC1 3.9 12295 1144 235 54 97 489 11.3 21.1 0.537 3.59 3.0 4.6 4.0 3.5 3.6 4.3 3.4 2.9 4.7 2.9 5 1CC1 - 1CC2 5.9 13041 1191 240 61 102 512 11.6 21.7 0.558 3.68 4.8 7.3 6.2 6.0 5.8 6.8 5.3 4.5 7.4 4.5 Pre-aeration 5 1CC1 - 1CC3 7.8 13484 1216 242 64 104 523 11.8 22.0 0.567 3.71 6.5 9.8 8.2 8.3 7.7 9.1 7.1 6.0 9.9 10.1

(SO2) and no 5 1CC1 - 1CC4 10.4 13992 1243 243 66 104 529 12.0 22.5 0.577 3.77 9.0 13.3 11.0 11.4 10.4 12.3 9.6 8.1 13.4 13.7 0.72

reagents. Dec-12 20 1CT (Average) 72.2 16415 944 228 60 105 442 13.11 29.7 0.434 2.80 91.0 86.7 89.0 88.6 89.6 87.7 90.4 91.9 86.6 86.3 Calc Head 16162 975 230 60 104 451 13.0 29.0 0.449 2.90

Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Baseline 8-12 1.16 1.18 0.99 1.18 1.09 Baseline 10-2 2.30 1.29 0.96 1.51 1.15 Z10-10-3 2.48 1.25 0.96 1.55 1.16 Z10-10-5 1.07 0.72 0.87 0.84 0.92

69

5.7. Pre-aeration Screening Tests – Rougher Concentrates

Since a significant improvement in separation was achieved between copper and bismuth in testwork on the rougher concentrate additional tests around the use of pre-aeration were completed. These tests were designed to determine if pre-aeration with air or SO2 alone were the sole contributors for the copper and bismuth separation. Table 5.11 summarizes the metallurgical results obtain from both tests as well as the most recent baseline test.

As shown through the data in Table 5.11 the rejection ratios for pre-aeration tests, air and SO2, were both lower than the baseline tests. The Pre-aeration test with air, Test Z10-10-3, had only a slightly lower rejection ratio than the baseline test. However, the rejection ratio in the pre-aeration test with SO2 was nearly half that of the baseline test. In both cases pre-aeration with air or SO2 was shown not be the sole contributor in the separation of copper and bismuth in the rougher concentrate. Based on the results presented reagents needs to be added in order to float both the copper and depress the bismuth after pre-aeration is competed on the rougher concentrate. Presented in Table 5.9 in Section 5.6, Depression of Bismuth Minerals – Rougher Concentrate, reagents were also used without pre-aeration prior to flotation and rejections ratios were less than those achieved in Test Z10-10-4, leading to the conclusion that the combination of pre-aeration followed by reagent additions for copper and bismuth need to be part of the flotation test conditions in order for the copper bismuth separation to be successful.

5.8. Summary of Successful Separation Tests – Final and Rougher Concentrates

Since it was shown that in both samples, final and rougher concentrates, the same set of conditions yielded the best separation results it was important to summarize the data and compare them based on all the penalty elements. Table 5.12 provides a summary with baseline and successful Cu/Bi separation test results for the final and rougher concentrates.

In Table 5.12 the rejection ratios are calculated for each element based on the recovery of copper divided by the recovery of the specific penalty element. The data shows that in both cases that pre- aeration with SO2 followed by the adjustment to pH >11.0, and the use of Cytec 3894/7262, that all penalty element rejection ratios are higher compared to baseline tests completed on each product. As stated previously the penalty elements from the ore at this particular deposit generally seem to behave in a similar manner, meaning if a single penalty element can be depressed then the whole suite will tend to follow the same trend.

5.9. Confirmation Test – Rougher Concentrate

A successful separation set of conditions had been found to remove a significant portion of the copper from the rougher concentrate leaving the bismuth behind. An additional test was completed on the rougher concentrate to validate the separation conditions that have been discussed previously.

70

Table 5.12 Metallurgical Results for Final and Rougher Concentrate Successful Pre-aeration Flotation Tests

Ret. Test Assay Distribution, % Reject Test Test % Regrind P80 Time Ratio Conditions Z10 Date mins. Solids Product wt. As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Final Concentrate Assayed Head 100 1186 633 153 36 128 529 16.55 20.63 0.263 1.15 8-13 n/a 59 5 8.0 1CC1 32.6 183 233 56 7 21 194 7.5 9.7 0.107 0.374 6.3 11.5 12.0 9.1 11.5 12.5 14.0 14.2 14.6 14.2 5 1CC1 - 1CC2 47.5 264 340 91 14 34 299 13.1 14.6 0.147 0.614 13.2 24.5 28.6 25.7 27.0 28.1 35.4 30.9 29.3 30.9 Baseline Test 5 1CC1 - 1CC3 54.9 319 402 111 17 38 344 14.9 16.2 0.165 0.746 18.4 33.5 40.1 37.2 35.0 37.4 46.5 39.8 38.1 39.2 Final 5 1CC1 - 1CC4 61.8 398 459 127 20 41 380 16.4 17.8 0.182 0.848 25.8 43.0 51.9 48.3 42.6 46.5 57.5 49.0 47.1 50.1 1.34

Concentrate 20 10-Oct-12 1CT (Average) 38.2 1856 980 191 34 90 709 19.51 29.9 0.330 1.37 74.2 57.0 48.1 51.7 57.4 53.5 42.5 51.0 52.9 49.9 Calc Head 955 658 152 25 60 506 17.6 22.4 0.239 1.05 8-5 n/a 59 5 10.0 1CC1 31.3 377 197 159 31 31 316 17.2 21.1 0.079 0.910 12.1 10.0 30.3 20.1 16.8 20.0 34.8 30.4 10.3 30.4 Pre-aeration 5 1CC1 - 1CC2 33.8 385 202 155 30 30 306 16.4 20.2 0.082 0.895 13.4 11.1 31.9 21.1 17.4 20.9 35.8 31.6 11.5 31.6

with SO2. 5 1CC1 - 1CC3 36.0 397 209 151 30 29 299 15.7 19.5 0.085 0.878 14.7 12.2 33.1 22.6 17.8 21.7 36.5 32.5 12.8 30.0 Final Concentrate Cytec 3894 5 1CC1 - 1CC4 37.8 409 214 148 30 28 293 15.1 19.0 0.087 0.862 15.9 13.1 33.9 23.7 18.1 22.3 37.0 33.2 13.7 30.9 2.82

and 7262. pH 10-Oct 20 11.0 1CT (Average) 62.2 1313 864 174 59 76 616 15.7 23.3 0.333 1.17 84.1 86.9 66.1 76.3 81.9 77.7 63.0 66.8 86.3 69.1 Calc Head 971 618 164 48 58 494 15.5 21.7 0.240 1.05 Rougher Concentrate Assayed Head 100 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 10-2 n/a 62 5 9.0 1CC1 43.0 6308 705 389 38 120 389 18.9 37.0 0.329 3.25 16.4 32.7 59.3 27.5 42.1 36.1 60.1 50.3 33.5 50.3 5 1CC1 - 1CC2 54.5 7437 803 376 43 125 413 17.5 36.9 0.377 3.23 24.5 47.2 72.6 39.3 55.5 48.6 70.8 63.7 48.7 63.7 Baseline Test 5 1CC1 - 1CC3 61.6 8345 862 365 47 129 428 16.9 36.8 0.399 3.24 31.1 57.3 79.7 48.4 65.0 56.9 77.0 71.7 58.2 62.5 Rougher 5 1CC1 - 1CC4 65.2 8821 886 360 48 131 434 16.6 36.7 0.407 3.26 34.8 62.3 83.2 52.9 69.4 61.1 80.1 75.7 62.7 66.6 1.29 Concentate 20 20-Dec-12 1CT (Average) 34.8 30919 1007 137 81 108 518 7.74 22.1 0.450 3.07 65.2 37.7 16.8 47.1 30.6 38.9 19.9 24.3 37.3 33.4 Calc Head 16519 928 282 59 123 463 13.5 31.6 0.422 3.20 10-4 n/a 62 5 8.0 1CC1 14.1 3173 331 173 10 85 395 28.2 27.8 0.142 1.84 2.8 4.9 10.6 2.3 11.2 12.7 30.2 13.4 4.7 13.4 5 1CC1 - 1CC2 24.1 3763 413 251 16 87 416 26.9 27.5 0.162 2.52 5.7 10.5 26.5 6.3 19.7 22.9 49.5 22.8 9.2 22.8 Pre-aeration 5 1CC1 - 1CC3 26.3 4348 463 258 19 88 425 26.1 27.3 0.188 2.66 7.2 12.8 29.7 8.4 21.8 25.5 52.4 24.7 11.6 23.7 with SO2 and

Rougher Concentrate Rougher 5 1CC1 - 1CC4 27.8 4867 504 257 22 89 432 25.4 27.1 0.207 2.72 8.5 14.7 31.3 10.1 23.3 27.4 53.8 25.9 13.5 25.7 3.66 Cytec 7262. 20-Dec 20 3894 for Cu. 1CT (Average) 72.2 20138 1129 218 75 114 440 8.39 29.9 0.514 3.03 91.5 85.3 68.7 89.9 76.7 72.6 46.2 74.1 86.5 74.3 Calc Head 15892 955 229 60 107 438 13.1 29.1 0.428 2.95

Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Baseline Final Concentrate 2.23 1.34 1.11 1.19 1.35 Z10-8-5 (Final Concentrate) 2.33 2.82 1.09 1.56 2.04 Baseline Rougher Concentrate 2.30 1.29 0.96 1.51 1.15 Z10-10-4 (Rougher Concentrate) 6.33 3.66 1.72 5.33 2.31

71

Table 5.13 summarizes the results for the successful tests that were presented in Table 5.12, but with the addition of the confirmation Test Z10-10-7 that utilized the same parameters as discussed in Section 7.8. It is important to note that Test Z10-10-7 required that the last remaining sample of rougher concentrate be split out prior to flotation testing. At this time the rougher concentrate had been received from operations approximately 2 years prior. Results for Test Z10-10-7 were similar to those shown in Test Z10-10-4 using the same set of conditions, but at slightly higher percent solids in Test Z10-10-7 compared to Test Z10-10-4. The rejection ratios for Test Z10-10-7 were slightly lower than those seen in the previous test, but a significant separation still occurred between copper and bismuth. Based on the confirmation test a assumption can be made that the percent solids during flotation may play a minor role in the separation between copper and bismuth. The density in the confirmation test was 16 % solids while the Test Z10-10-4 was 10 % which could account for the lower rejections ratios is the confirmation test. As was the case for the testwork completed on the final concentrate, aging of the feed sample may also have played a role and impacted the results.

72

Table 5.13: Metallurgical Results for Confirmation of Successful Test Conditions on Rougher Concentrate

Test Assay Distribution, % Reject Test Test Ret. % Z10 Product wt. As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn Ratio Conditions Date Time Solids % ppm ppm ppm ppm ppm gpt % % % % Cu/Bi Final Concentrate Assayed Head 100 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 8-13 5 9.0 1CC1 32.6 183 233 56 7 21 194 7.5 9.7 0.107 0.374 6.3 11.5 12.0 9.1 11.5 12.5 14.0 14.2 14.6 14.2 5 1CC1 - 1CC2 47.5 264 340 91 14 34 299 13.1 14.6 0.147 0.614 13.2 24.5 28.6 25.7 27.0 28.1 35.4 30.9 29.3 30.9 Baseline Test 5 1CC1 - 1CC3 54.9 319 402 111 17 38 344 14.9 16.2 0.165 0.746 18.4 33.5 40.1 37.2 35.0 37.4 46.5 39.8 38.1 39.2 Final 5 1CC1 - 1CC4 61.8 398 459 127 20 41 380 16.4 17.8 0.182 0.848 25.8 43.0 51.9 48.3 42.6 46.5 57.5 49.0 47.1 50.1 1.34

Concentrate 20 10-Oct-12 1CT (Average) 38.2 1856 980 191 34 90 709 19.51 29.9 0.330 1.37 74.2 57.0 48.1 51.7 57.4 53.5 42.5 51.0 52.9 49.9 Calc Head 955 658 152 25 60 506 17.6 22.4 0.239 1.05 8-5 5 10.0 1CC1 31.3 377 197 159 31 31 316 17.2 21.1 0.079 0.910 12.1 10.0 30.3 20.1 16.8 20.0 34.8 30.4 10.3 30.4 Pre-aeration 5 1CC1 - 1CC2 33.8 385 202 155 30 30 306 16.4 20.2 0.082 0.895 13.4 11.1 31.9 21.1 17.4 20.9 35.8 31.6 11.5 31.6

with SO2. 5 1CC1 - 1CC3 36.0 397 209 151 30 29 299 15.7 19.5 0.085 0.878 14.7 12.2 33.1 22.6 17.8 21.7 36.5 32.5 12.8 30.0 Final Concentrate Cytec 3894 5 1CC1 - 1CC4 37.8 409 214 148 30 28 293 15.1 19.0 0.087 0.862 15.9 13.1 33.9 23.7 18.1 22.3 37.0 33.2 13.7 30.9 2.82

and 7262. pH 10-Oct 20 11.0 1CT (Average) 62.2 1313 864 174 59 76 616 15.7 23.3 0.333 1.17 84.1 86.9 66.1 76.3 81.9 77.7 63.0 66.8 86.3 69.1 Calc Head 971 618 164 48 58 494 15.5 21.7 0.240 1.05

Rougher Concentrate Assayed Head 100 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 10-2 5 9.0 1CC1 43.0 6308 705 389 38 120 389 18.9 37.0 0.329 3.25 16.4 32.7 59.3 27.5 42.1 36.1 60.1 50.3 33.5 50.3 5 1CC1 - 1CC2 54.5 7437 803 376 43 125 413 17.5 36.9 0.377 3.23 24.5 47.2 72.6 39.3 55.5 48.6 70.8 63.7 48.7 63.7 Baseline Test 5 1CC1 - 1CC3 61.6 8345 862 365 47 129 428 16.9 36.8 0.399 3.24 31.1 57.3 79.7 48.4 65.0 56.9 77.0 71.7 58.2 62.5 Rougher 5 1CC1 - 1CC4 65.2 8821 886 360 48 131 434 16.6 36.7 0.407 3.26 34.8 62.3 83.2 52.9 69.4 61.1 80.1 75.7 62.7 66.6 1.29 Concentate 20 20-Dec-12 1CT (Average) 34.8 30919 1007 137 81 108 518 7.74 22.1 0.450 3.07 65.2 37.7 16.8 47.1 30.6 38.9 19.9 24.3 37.3 33.4 Calc Head 16519 928 282 59 123 463 13.5 31.6 0.422 3.20 10-4 5 8.0 1CC1 14.1 3173 331 173 10 85 395 28.2 27.8 0.142 1.84 2.8 4.9 10.6 2.3 11.2 12.7 30.2 13.4 4.7 13.4 5 1CC1 - 1CC2 24.1 3763 413 251 16 87 416 26.9 27.5 0.162 2.52 5.7 10.5 26.5 6.3 19.7 22.9 49.5 22.8 9.2 22.8 Pre-aeration 5 1CC1 - 1CC3 26.3 4348 463 258 19 88 425 26.1 27.3 0.188 2.66 7.2 12.8 29.7 8.4 21.8 25.5 52.4 24.7 11.6 23.7 with SO2 and 5 1CC1 - 1CC4 27.8 4867 504 257 22 89 432 25.4 27.1 0.207 2.72 8.5 14.7 31.3 10.1 23.3 27.4 53.8 25.9 13.5 25.7 3.66 Cytec 7262. 20-Dec 20 3894 for Cu. 1CT (Average) 72.2 20138 1129 218 75 114 440 8.39 29.9 0.514 3.03 91.5 85.3 68.7 89.9 76.7 72.6 46.2 74.1 86.5 74.3

Calc Head 15892 955 229 60 107 438 13.1 29.1 0.428 2.95 Rougher Concentrate Rougher Confirmation Test - Rougher Concentrate 10-7 5 8.0 1CC1 1.6 12054 1053 223 76 71 15 14.2 24.5 0.466 2.71 1.1 1.7 1.5 1.5 1.6 1.7 1.5 1.2 1.7 1.2 5 1CC1 - 1CC2 5.9 11134 954 203 72 62 15 18.1 26.9 0.419 2.53 3.9 5.7 5.1 5.4 5.3 6.5 7.4 4.9 5.7 4.9 Pre-aeration 5 1CC1 - 1CC3 21.6 6405 598 294 46 50 15 25.5 27.8 0.230 2.60 8.3 13.1 27.2 12.6 15.7 23.1 38.3 18.4 11.5 23.0 with SO2 and 5 1CC1 - 1CC4 27.6 7800 720 288 53 54 15 23.6 27.6 0.289 2.73 12.9 20.2 33.9 18.6 21.3 30.4 45.2 23.3 18.4 30.8 2.24 Cytec 7262. Jan-13 20 3894 for Cu. 1CT (Average) 72.4 20166 1082 213 89 76 13 10.9 34.9 0.488 2.33 87.1 79.8 66.1 81.4 78.7 69.6 54.8 76.7 81.6 69.2 Calc Head 16753 982 234 79 70 14 14.4 32.9 0.433 2.44

Rejection Ratios by Element Cu/As Cu/Bi Cu/Cd Cu/Sb Cu/Se Baseline Final Concentrate 2.23 1.34 1.11 1.19 1.35 Z10-8-5 (Final Concentrate) 2.33 2.82 1.09 1.56 2.04 Baseline Rougher Concentrate 2.30 1.29 0.96 1.51 1.15 Z10-10-4 (Rougher Concentrate) 6.33 3.66 1.72 5.33 2.31 Z10-10-7 (Rougher Concentrate) 3.50 2.24 1.33 2.43 2.12

73

CHAPTER 6 CONCLUSIONS AND RECOMMENDATIONS

A separation of copper minerals from bismuth minerals can involve the use of many different mineral processing (mainly flotation) and hydrometallurgical processes. A review of the past flotation literature offered some suggestions on where to begin regarding flotation parameters to investigate, but the majority of the literature discussed the separation of arsenic with very limited literature available for the remaining major penalty elements such as bismuth. In general, with a few exceptions the penalty elements in the case of this research program seemed to behave similarly based on the trends seen in the testwork completed and presented. The objective of this research program was to identify a chemistry that may be applicable specifically to a bismuth flotation separation, as very little is known of bismuth flotation separations in the literature. This research program identified a set of flotation conditions that showed great promise in achieving the objective of making a bismuth separation, although further work will be needed to improve copper recoveries, as well as optimize the various reagents/conditions identified.

Penalty element mineralogy was of great importance in the testwork that was conducted. Extensive mineralogy was completed on five samples received from a copper concentrator. Due to the penalty element levels the main focus and mineralogical characterization focused on bismuth. Several bismuth minerals were observed in the samples, bismuthinite, tetradymite, AgPbBiS3, and PbBiSbAgS.

The majority of bismuth observed in the samples occurred within the unidentified mineral AgPbBiS3. Mineralogical associations between copper and bismuth were shown to be minimal, but some associations did exist which could be investigated further. Mineral liberation analysis showed that a vast majority of the bismuth existed at fine mineral grain sizes in the samples tested, < 25 microns. This potentially identified a problem since the separation technique that was to be applied was flotation. Although flotation at fine particle sizes can be achieved it can pose some difficulty in the separation.

Much of the testwork completed throughout this thesis offered insight on flotation parameters that showed the potential of achieving a separation of penalty elements from copper, but most results were not significant enough to investigate further. However, a single set of conditions was found to be promising on both the final and rougher concentrate tested. The process included pre-aeration of the flotation feed with SO2. The SO2 is believed to offer surface conditioning of minerals in the samples. Surface preparation is thought to include stripping of reagents present in the sample prior to testwork.

SO2 also seems to offer the ability to depress the vast majority of minerals in the samples tested allowing them to become re-activated/depressed with the reagents used in the testwork. The pH was adjusted to > 11.0 based on literature surrounding the rejection of arsenopyrite at alkaline pH and to reject

74 pyrite/gangue to limit the dilution of the copper concentrate. Cytec 7262 was known as a flotation depressant for lead, but results suggest it also offers penalty element depress capabilities. A portion of the bismuth was shown through mineralogy to be associated with the mineral AgPbBiS3, however the remaining penalty elements were also reduced in both concentrates. Although not a single solution was found to result in complete penalty element depress, a series of flotation modifications can be made to reduce penalty elements while achieving decent copper grade and recovery.

It is important to point out the reproducibility of the baseline tests for both the final and rougher concentrates. In both cases the repeat baseline test series for both products reproduced well generating similar results. Although some very minor differences did occur this possibly could be accounted for through the aging of the sample since the repeat tests were completed many months from the original baseline tests. With that being said the test conditions that were shown to offer a significant separation worked for both products and similar results were obtained when compared back to their respective baseline tests. This is also an important pre-caution for laboratory testwork as to how long a previously processed sample remains viable for further laboratory testwork. It further illustrates the importance of ensuring that baseline tests be completed routinely when a new test series is initiated to be able to identify if sample degradation may be occurring, so that testwork results can be reviewed and interpreted within this context. The accountabilities between the majority of the baseline tests with either the final concentrate or rougher concentrate were excellent for copper and lead, which gives confidence in the results. The other penalty elements such as cadmium, selenium and antimony varied, most likely due to their low levels for detection through chemical analysis. Arsenic in most cases as shown previously balanced well in all tests conducted.

Additional information was discovered around some of the reagents tested in this program. As shown in the previous results section, the Nalco reagents were seen to be excellent copper collectors. Although the original information received from Nalco indicated that these reagents were selective for copper while depressing penalty elements. No separation of copper from bismuth was observed to have taken place based on results produced from the tests utilizing the Nalco reagents. Perhaps a point of future work would be to utilize the Nalco reagents in conjunction with Cytec 7262 which has been shown to be a penalty element depression reagent through the tests completed in this thesis when used in combination with SO2 for pre-aeration.

Future testwork considering optimization of the identified bismuth flotation separation parameters should be pursued. Investigations should explore if reagent dosages are sufficient or if additional separations could be made with additional 7262. Additional conditioning time with SO2 should also be investigated to ensure adequate time was given or if the 30 minutes utilized in this testwork could be reduced or would need to be expanded. Other testwork that would be interesting to consider in future

75 work, would be to detailed mineralogy on the products from the bismuth flotation separation to evaluate which bismuth minerals (all or are some preferentially impacted) are impacted by the flotation conditions identified.

76

References Cited

Bruckard, W.J., Kyriakidis, I., Woodcock, J.T. "The flotation of metallic arsenic as a function of pH and pulp potential - A single mineral study." International Journal of Mineral Processing 84 (2007): 25- 32.

Bulatovic, S. M. Hanbook of Flotation Reagents. Amsterdam: Elsevier B.V. , 2007.

Cytec Mineral Processing Chemicals Technology. "Innovations in Flotation Reagents." Presentation. Englewood, CO: Cytec Industries, January 2013.

Dai, Z., Garritsen, J., Wells, P F., Xu, M. "Arsenic Rejection in the Flotation of Garson Ni-Cu Ore." Centenary of Flotation Symposium. Brisbane, 2005. 939-946.

Doerr, D.L., D. Lopez, R. Kappes, and T.J. Drake. "APSEX CORP White Papers." APSEX CORP. 2009. http://aspexcorp.com/Portals/0/whitepapers/alternative%20-electron%20beam%20technology.pdf (accessed January 15, 2012).

F, Reid A, and Wittenberg J C. "Technical Note - Rapid Production of High Quality Polished Sections for Automated Image Analysis of Minerals." AUSIMM, 1984: 5.

Fabiano Capponi, Elves Matiolo, Denise G. Nunes, Jorge Rubio. "ADVANCES IN FLOTATION OF MINERALS FINES." Research Findings. : Codelco-División, 1999.

Fornasiero, D., Fullston, D., C. Li., Ralston, J. "Separation of enargite and tennantite from non-arsenic copper sulfide minerals by selective oxidation or dissolution." Internation Journal of Mineral Processing 61 (2001): 109-119.

Gathje, John, interview by Zach Zanetell. Penalty Element Depression (January 15, 2010).

Kappes, R., Brosnahan, D., Gathje, J. "Characterization of Copper Flotation Products Utilizing The JKMRC/FEI Mineral Liberation Analyzer." SME Annual Meeting. Denver: Society of Mining Metallurgy and Exploration, 2007. 07-061.

Kelly, E.G., and D.J. Spottiswood. Introduction to Mineral Processing. Auckland: John Wiley & Sons, 1989.

Kennedy, B.A. 2nd Edition. Society of Mining Metallurgy and Exploration, 1990.

77

Khmelnitskaya, O D., Beskrovnaya, V P. "A Technology of Precious Metals Recovery from Gold-Bismuth Ores." World Gold Conference. Cairns: World Gold, 2007. 22-24.

Kouzmanov, K., Bogdanov, K., Ramboz, C. "Te- and Bi-bearing assemblages in the Elshitsa Radka epitherman deposite, Central Srednogorie, Bulgaria: Mineralogy and genetical features." Geochemistry, Mineralogy and Petrology 43 (2005): 108-112.

Laney, Debbie, interview by Zach Zanetell. Senior Technical Consultant (January 15, 2010).

Ma, X., Bruckard, W.J. "Rejection of Arsenic Minerals in Sulfide Flotation - A Literature Review." International Journal of Mineral Processing, 2009: 89-94.

Maes, Charles, interview by Zach Zanetell. Senior Technical Expert (January 15, 2011).

Mular, A. L., D. N. Halbe, and D. J. Barratt. Mineral Processing Plant Design, Practice and Control. Littleton: Society for Mining Metallurgy and Exploration, 2002.

Newcombie, L. A. "Differential Flotation at Warrego Concentrator." The Aus.I.M.M Conference. Darwin: Aus.I.M.M, 1984. 291-298.

Newmont Mining Corporation. "Penalty Element Contract Between Xstrata and A Copper Smelter." Elko: Newmont Mining, 2011.

Rao, S. R., and J. Leja. Surface Chemistry of Froth Flotation. New York: Kluwer Academic/Plenum Publishers, 2004.

Smith, G. Sampling in the mineral and metallurgical processing industries. London: Institution of Mining and Metallurgy, 1973.

Steinhart, Terry L. Partilce Size Testing Methodology. 2010. http://www.swinefeedefficiency.com/factsheets/IPIC25c%20SFE%20Particle%20Size%20Testing %20Methodology.pdf (accessed March 24, 2013).

Subramanian, K N., Connelly, D E G., Wong, K Y. "Separation of Pyrtie and Arsenopyrite in a Gold Sulfide Concentrate." Centenary of Flotation Symposium. Bisbane, 2005. 1045-1052.

Thomas, Willard. Mining Chemicals Handbook. USA: Cytec Industries, Inc., 2010.

78

Zhengzhou ZY Machinery CO.,LTD. Flotation Machines. n.d. www.zzywzg.com (accessed April 3, 2013).

79

APPENDIX A

80

Table A-1: Particle Size Analysis Final Concentrate

Particle Size Analysis

Test No: Z10-4-1

Lot # 110322-1

Sample: Final Conc

Grind: Time, minAs Received % Solids Mill

Procedure: A representative split of the sample was wet screened at 500 mesh and the plus fraction was dried. It was then dry screened by Ro-Tapping at the sizes indicated.

Results:

Product Weight Tyler µm Retained Cumulative Mesh g % Retained Passing

65 212 0.0 0.0 0.0 100.0 100 150 0.0 0.0 0.0 100.0 150 105 3.6 3.3 3.3 96.7 200 75 9.4 8.6 11.9 88.1 270 53 11.9 10.9 22.8 77.2 325 45 5.2 4.8 27.6 72.4 400 38 4.4 4.0 31.6 68.4 500 25 12.6 11.6 43.2 56.8 Pan -25 61.7 56.8 100.0 Total 108.8 100.0

P80 = 59 µm

Cumulative Weight % Passing as a Function of Particle Size

100.0

90.0

80.0

70.0

60.0

50.0

40.0

Data Points Cumulative % Wt Passing 30.0

20.0 0 20 40 60 80 100 120 140 160 180 200 220 240 260 Particle Size, µm

81

A-2: Particle Size Analysis Rougher Concentrate

Particle Size Analysis

Test No: Z10-4-2

Lot # 110322-2

Sample: Ro. Conc.

Grind: Time, minAs Received % Solids Mill

Procedure: A representative split of the sample was wet screened at 500 mesh and the plus fraction was dried. It was then dry screened by Ro-Tapping at the sizes indicated.

Results:

Product Weight Tyler µm Retained Cumulative Mesh g % Retained Passing

65 212 1.0 0.8 0.8 99.2 100 150 2.1 1.6 2.4 97.6 150 105 4.6 3.5 5.9 94.1 200 75 10.3 7.8 13.7 86.3 270 53 13.8 10.5 24.2 75.8 325 45 5.9 4.5 28.7 71.3 400 38 5.6 4.3 33.0 67.0 500 25 14.1 10.7 43.7 56.3 Pan -25 74.1 56.3 100.0 Total 131.5 100.0

P80 = 62 µm

Cumulative Weight % Passing as a Function of Particle Size

100.0

90.0

80.0

70.0

60.0

50.0

40.0

Data Points Cumulative % Wt Passing 30.0

20.0 0 20 40 60 80 100 120 140 160 180 200 220 240 260 Particle Size, µm

82

Table A-3: Particle Size Analysis Rougher Scavenger Concentrate

Particle Size Analysis

Test No: Z10-4-3

Lot # 110322-3

Sample: Ro Scav Con

Grind: Time, minAs Received % Solids Mill

Procedure: A representative split of the sample was wet screened at 500 mesh and the plus fraction was dried. It was then dry screened by Ro-Tapping at the sizes indicated.

Results:

Product Weight Tyler µm Retained Cumulative Mesh g % Retained Passing

65 212 0.0 0.0 0.0 100.0 100 150 2.0 2.3 2.3 97.7 150 105 3.5 4.1 6.4 93.6 200 75 8.5 9.8 16.2 83.8 270 53 9.4 10.9 27.1 72.9 325 45 4.0 4.6 31.7 68.3 400 38 2.4 2.8 34.5 65.5 500 25 8.1 9.4 43.9 56.1 Pan -25 48.4 56.1 100.0 Total 86.3 100.0

P80 = 67 µm

Cumulative Weight % Passing as a Function of Particle Size

100.0

90.0

80.0

70.0

60.0

50.0

40.0

Data Points Cumulative % Wt Passing 30.0

20.0 0 20 40 60 80 100 120 140 160 180 200 220 240 260 Particle Size, µm

83

Table A-4: Particle Size Analysis Rougher Feed

Particle Size Analysis

Test No: Z10-4-4

Lot # 110322-5

Sample: Ro Feed

Grind: Time, minAs Received % Solids Mill

Procedure: A representative split of the sample was wet screened at 500 mesh and the plus fraction was dried. It was then dry screened by Ro-Tapping at the sizes indicated.

Results:

Product Weight Tyler µm Retained Cumulative Mesh g % Retained Passing

65 212 5.4 6.5 6.5 93.5 100 150 8.6 10.4 16.9 83.1 150 105 7.6 9.2 26.1 73.9 200 75 8.2 9.9 36.0 64.0 270 53 7.1 8.6 44.6 55.4 325 45 2.6 3.1 47.7 52.3 400 38 1.0 1.2 48.9 51.1 500 25 6.8 8.2 57.1 42.9 Pan -25 35.6 42.9 100.0 Total 82.9 100.0

P80 = 135 µm

Cumulative Weight % Passing as a Function of Particle Size

100.0

90.0

80.0

70.0

60.0

50.0

40.0

Data Points Cumulative % Wt Passing 30.0

20.0 0 20 40 60 80 100 120 140 160 180 200 220 240 260 Particle Size, µm

84

Table A-5: Particle Size Analysis Rougher Tails

Particle Size Analysis

Test No: Z10-4-5

Lot # 110322-4

Sample: Ro Tails

Grind: Time, minAs Received % Solids Mill

Procedure: A representative split of the sample was wet screened at 500 mesh and the plus fraction was dried. It was then dry screened by Ro-Tapping at the sizes indicated.

Results:

Product Weight Tyler µm Retained Cumulative Mesh g % Retained Passing

65 212 24.7 11.0 11.0 89.0 100 150 29.2 13.0 24.0 76.0 150 105 21.9 9.7 33.7 66.3 200 75 21.2 9.4 43.1 56.9 270 53 17.4 7.7 50.8 49.2 325 45 7.1 3.2 54.0 46.0 400 38 4.8 2.1 56.1 43.9 500 25 13.3 5.9 62.0 38.0 Pan -25 85.7 38.0 100.0 Total 225.3 100.0

P80 = 169 µm

Cumulative Weight % Passing as a Function of Particle Size

100.0

90.0

80.0

70.0

60.0

50.0

40.0

Data Points Cumulative % Wt Passing 30.0

20.0 0 20 40 60 80 100 120 140 160 180 200 220 240 260 Particle Size, µm

85

APPENDIX B

86

Table B-1: Cyclosizer Data Final Concentrate

Cyclosizer Data Sheet

Sample: Z10-6-1 Lot #: 110322

Starting weight 14.4840 grams Temperature 20.0 ºC Factor = 1.00 Flow 206 mm Factor = 0.93 Specific Gravity 2.70 Factor = 0.98 Time 20 minutes Factor = 0.93 Overall Factor = 0.85

Weight Size, µ grams % Cum % Cum % Cyclone Stand Fact'd Retain Pass

1 43.7 37.0 3.9408 27.21 27.21 72.79 2 31.8 26.9 0.2810 1.94 29.15 70.85 3 22.1 18.7 0.4223 2.92 32.07 67.93 4 15.4 13.0 0.8193 5.66 37.73 62.27 5 12.4 10.5 1.0152 7.01 44.74 55.26

Total 6.4786

P80 = 74 µm Percent Passing 10µm = 54

Particle Size Distribution 100.0 90.0 80.0 70.0 60.0 50.0 40.0 30.0

Cumulative % Passing 20.0 10.0 0.0 0 5 10 15 20 25 30 35 40

Passing Size, µm

87

Table B-2: Cyclosizer Data Final Concentrate 5 Minute Regrind

Cyclosizer Data Sheet

Sample: Z10-6-2 Lot #: 110322

Starting weight 28.2979 grams Temperature 20.0 ºC Factor = 1.00 Flow 206 mm Factor = 0.93 Specific Gravity 2.70 Factor = 0.98 Time 20 minutes Factor = 0.93 Overall Factor = 0.85

Weight Size, µ grams % Cum % Cum % Cyclone Stand Fact'd Retain Pass

1 43.7 37.0 2.1464 7.59 7.59 92.41 2 31.8 26.9 1.7014 6.01 13.60 86.40 3 22.1 18.7 2.3908 8.45 22.05 77.95 4 15.4 13.0 3.2066 11.33 33.38 66.62 5 12.4 10.5 3.1953 11.29 44.67 55.33

Total 12.6405

P80 = 21 µm Percent Passing 10µm = 53

Particle Size Distribution 100.0 90.0 80.0 70.0 60.0 50.0 40.0 30.0

Cumulative % Passing 20.0 10.0 0.0 0 5 10 15 20 25 30 35 40

Passing Size, µm

88

Table B-3: Cyclosizer Data Final Concentrate 10.0 Minute Regrind

Cyclosizer Data Sheet

Sample: Z10-6-3 Lot #: 110322

Starting weight 39.7333 grams Temperature 20.0 ºC Factor = 1.00 Flow 206 mm Factor = 0.93 Specific Gravity 2.70 Factor = 0.98 Time 20 minutes Factor = 0.93 Overall Factor = 0.85

Weight Size, µ grams % Cum % Cum % Cyclone Stand Fact'd Retain Pass

1 43.7 37.0 0.5771 1.45 1.45 98.55 2 31.8 26.9 1.2374 3.11 4.56 95.44 3 22.1 18.7 3.2876 8.27 12.83 87.17 4 15.4 13.0 6.1022 15.36 28.19 71.81 5 12.4 10.5 5.8354 14.69 42.88 57.12

Total 17.0397

P80 = 16 µm Percent Passing 10µm = 54

Particle Size Distribution 100.0 90.0 80.0 70.0 60.0 50.0 40.0 30.0

Cumulative % Passing 20.0 10.0 0.0 0 5 10 15 20 25 30 35 40

Passing Size, µm

89

APPENDIX C

90

Table C-1: Copper Mineral Liberation by Free Surface for Final Concentrate

Enargite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80=34 No. of Particles 31344 13 43 10 6 4 2 24 Mean Phase ECD (µm) 5.72 7.90 7.54 14.15 13.24 66.42 15.39 Mean Phase Max Span (µm) 9.26 13.04 11.94 21.66 21.45 82.90 21.86 Particle Distribution (%) 99.46 0.16 0.23 0.04 0.02 0.00 0.01 0.07 Mean Density 3.75 4.27 3.64 4.26 4.29 4.37 1.02 4.40 Distribution of Mineral (%) 1.22 12.05 9.55 9.42 2.71 9.61 55.45 Cum. Distn. of Mineral (%) 100.00 98.78 86.73 77.18 67.77 65.06 55.45

Covellite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80=34 No. of Particles 30571 48 263 114 76 48 32 294 Mean Phase ECD (µm) 4.41 5.64 7.94 10.69 13.51 31.05 11.52 Mean Phase Max Span (µm) 7.04 8.88 13.59 18.21 22.00 48.63 16.41 Particle Distribution (%) 96.25 0.41 1.42 0.42 0.34 0.20 0.20 0.75 Mean Density 3.74 3.95 4.03 4.27 4.36 4.51 4.65 4.70 Distribution of Mineral (%) 0.16 4.67 6.53 11.49 10.03 13.28 53.84 Cum. Distn. of Mineral (%) 100.00 99.84 95.18 88.64 77.15 67.12 53.84

Chalcopyrite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80=34 No. of Particles 14346 139 1118 899 780 824 869 12471 Mean Phase ECD (µm) 4.94 8.23 12.09 20.02 28.78 44.22 18.94 Mean Phase Max Span (µm) 7.23 12.57 18.78 30.93 43.42 65.19 27.77 Particle Distribution (%) 36.30 0.65 5.62 4.43 3.31 3.43 4.11 42.15 Mean Density 3.38 3.83 3.57 3.73 3.95 4.09 4.15 4.19 Distribution of Mineral (%) 0.02 0.91 2.52 3.62 5.26 7.38 80.29 Cum. Distn. of Mineral (%) 100.00 99.98 99.07 96.55 92.93 87.67 80.29

Chalcocite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80=34 No. of Particles 31445 1 Mean Phase ECD (µm) 10.34 Mean Phase Max Span (µm) 12.79 Particle Distribution (%) 99.99 0.01 Mean Density 3.76 1.08 Distribution of Mineral (%) 100.00 Cum. Distn. of Mineral (%) 100.00

91

Table C-2: Copper Mineral Liberation by Free Surface for Rougher Concentrate

Enargite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80=34 No. of Particles 30182.00 19.00 87.00 28.00 23.00 16.00 14.00 73.00 Mean Phase ECD (µm) 4.14 6.42 7.99 9.72 16.81 34.77 12.40 Mean Phase Max Span (µm) 5.83 10.42 13.23 15.16 24.73 51.99 17.40 Particle Distribution (%) 96.52 1.38 1.68 0.08 0.05 0.08 0.05 0.16 Mean Density 4.33 4.30 4.24 4.43 4.31 4.33 4.38 4.40 Distribution of Mineral (%) 1.03 7.86 4.55 6.84 18.03 13.88 47.81 Cum. Distn. of Mineral (%) 100.00 98.97 91.10 86.56 79.72 61.69 47.81

Covellite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80=21 No. of Particles 29311.00 64.00 302.00 108.00 100.00 76.00 49.00 432.00 Mean Phase ECD (µm) 3.74 5.00 7.70 9.47 12.98 20.76 10.40 Mean Phase Max Span (µm) 5.43 7.70 13.11 17.41 24.02 34.27 14.58 Particle Distribution (%) 89.93 1.88 6.07 0.52 0.32 0.29 0.17 0.81 Mean Density 4.33 4.45 4.37 4.32 4.43 4.54 4.59 4.69 Distribution of Mineral (%) 0.60 9.09 8.00 9.44 11.66 9.01 52.20 Cum. Distn. of Mineral (%) 100.00 99.40 90.31 82.31 72.87 61.21 52.20

Chalcopyrite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80=63 No. of Particles 17427.00 122.00 803.00 750.00 701.00 758.00 826.00 9055.00 Mean Phase ECD (µm) 5.20 7.39 11.01 15.22 23.20 41.94 19.62 Mean Phase Max Span (µm) 7.35 11.42 17.02 23.38 34.86 61.99 28.68 Particle Distribution (%) 49.93 0.74 4.51 4.48 7.02 5.44 4.34 23.54 Mean Density 4.44 4.25 4.37 4.22 4.22 4.21 4.20 4.19 Distribution of Mineral (%) 0.06 0.81 3.70 10.18 10.87 10.84 63.53 Cum. Distn. of Mineral (%) 100.00 99.94 99.13 95.43 85.25 74.37 63.53

Chalcocite Free Surface of Particle (%) 0% (barren) 0% (not exposed) 0% < x <= 20% 20% < x <= 40% 40% < x <= 60% 60% < x <= 80% 80% < x < 100% 100%

P80=52 No. of Particles 30435.00 2.00 4.00 1.00 Mean Phase ECD (µm) 2.17 6.62 52.50 Mean Phase Max Span (µm) 2.94 10.02 71.20 Particle Distribution (%) 99.96 0.01 0.03 0.00 Mean Density 4.33 1.05 1.06 1.37 Distribution of Mineral (%) 0.30 13.09 86.61 Cum. Distn. of Mineral (%) 100.00 99.70 86.61 86.61 86.61 86.61 86.61

92

APPENDIX D

93

Table D-1: Flotation Data Sheet Z10-6-1

Composite: Plant Final Concentrate Date: 23-Jan-12 Z10 6-1 Plant Sample from Feb 10, 2009 Purpose: Baseline Test Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Penalty element separation baseline test series Regrind: RMA % minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell pH W22C H55 Stage(s) Grind Cond. Float mv wt (g) liters Regrind 0.0 7.92 140 Cond 5.0 7.95 65 1CC1 51 17 1.0 107 2.3 1CC2 17 17 2.0 8.21 151 160 2.3 1CC3 34 3.0 8.24 151 225 2.3 1CC4 34 5.0 8.19 225 270 2.3 1CC5 17 17 5.0 8.21 225 221 2.3

CT= 2407

Total ###### 51.000 0.0 5.0 16.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm opt % % % % 1CC1 34.73 14.94 32.46 209 248 124 16 42 252 12.3 14.9 0.125 0.747 3.1 6.1 9.9 7.2 8.5 7.3 10.6 10.1 7.7 10.1 1CC2 46.02 19.80 28.76 294 356 167 23 56 357 15.9 17.9 0.170 0.959 5.8 11.6 17.7 13.8 15.0 13.8 18.1 16.1 13.9 17.2 1CC3 56.51 24.31 25.12 789 609 224 37 80 535 22.4 24.7 0.255 1.30 19.2 24.5 29.2 27.2 26.4 25.4 31.2 27.3 25.7 28.6 1CC4 35.48 15.26 13.14 1760 739 217 40 80 577 20.2 27.8 0.271 1.25 27.0 18.6 17.7 18.5 16.6 17.2 17.7 19.3 17.1 17.3 1CC5 15.91 6.84 7.20 1992 732 185 39 86 587 17.3 27.5 0.241 1.08 13.7 8.3 6.8 8.1 8.0 7.8 6.8 8.6 6.8 6.7

1CTA 1615 990 187 44 104 781 14.4 20.9 0.375 1.21 1CTB 1679 183 44 95 770 14.5 22.1 0.360 1.15

1CT (Average) 43.83 18.85 1.82 1647 990 185 44.0 99.5 776 14.5 21.5 0.4 1.18 31.2 30.9 18.7 25.2 25.5 28.5 15.6 18.6 28.8 20.1

Calculated Heads 232.48 100.00 997 605 187 33 74 513 17.4 22.0 0.241 1.11 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 Accountabilities 84.0% 95.6% 122% 91.7% 57.5% 96.9% 105.2% 106.4% 91.6% 96.1%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm opt % % % % 1CC1 34.73 14.94 32.46 209 248 124 16 42 252 12.3 14.9 0.125 0.747 3.1 6.1 9.9 7.2 8.5 7.3 10.6 10.1 7.7 10.1 1CC1 - 1CC2 80.75 34.73 30.24 257 310 149 20 50 311.8 14.4 16.6 0.150 0.868 8.9 17.7 27.6 21.0 23.5 21.1 28.7 26.2 21.6 26.2 1CC1 - 1CC3 137.26 59.04 27.90 476 433 180 27 62 404 17.6 19.9 0.193 1.05 28.1 42.2 56.8 48.2 49.9 46.5 59.9 53.5 47.3 55.9 1CC1 - 1CC4 172.74 74.30 22.67 740 496 187 30 66 439 18.2 21.6 0.209 1.09 55.1 60.8 74.5 66.7 66.5 63.7 77.6 72.8 64.4 73.2 1CC1 - 1CC5 188.65 81.15 19.19 846 516 187 30 68 452 18.1 22.1 0.212 1.09 68.8 69.1 81.3 74.8 74.5 71.5 84.4 81.4 71.2 79.9

94

Table D-2: Flotation Data Sheet Z10-6-2

Composite: Plant Final Concentrate Date: 23-Jan-12 Z10 6-2 Plant Sample from Feb 10, 2009 Purpose: Baseline Test Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Penalty element separation baseline test series Regrind: RMA % 5.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell pH W22C H55 Stage(s) Grind Cond. Float mv wt (g) liters Regrind 5.0 8.21 96 Cond 5.0 8.21 50 1CC1 17 34 1.0 8.30 117 90 2.3 1CC2 34 2.0 8.27 144 149 2.3 1CC3 34 3.0 8.23 148 179 2.3 1CC4 17 34 5.0 8.14 167 251 2.3 1CC5 34 5.0 8.16 189 281 2.3

CT= 2359

Total 68.000 ###### 5.0 5.0 16.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm opt % % % % 1CC1 19.77 7.66 21.97 203 181 88 11 48 212 9.2 11.2 0.093 0.574 1.5 2.2 3.6 2.6 3.9 3.1 4.1 3.9 2.8 3.9 1CC2 31.61 12.25 21.21 231 233 133 15 53 300 14.7 15.7 0.122 0.806 2.7 4.6 8.8 5.6 6.9 7.1 10.5 8.8 5.9 8.7 1CC3 39.46 15.30 22.04 328 377 179 24 76 416 20.5 20.8 0.181 1.07 4.8 9.2 14.8 11.3 12.3 12.2 18.3 14.5 10.9 14.4 1CC4 39.88 15.46 15.89 597 598 209 37 98 494 22.3 23.3 0.266 1.19 8.8 14.8 17.5 17.6 16.1 14.7 20.1 16.4 16.2 16.2 1CC5 20.75 8.04 7.38 984 846 230 47 115 575 21.2 24.2 0.285 1.33 7.5 10.9 10.0 11.6 9.8 8.9 10.0 8.8 9.0 9.4

1CTA 1916 884 204 39 118 681 15.0 25.9 0.338 1.28 1CTB 1903 201 42 115 678 15.6 24.8 0.342 1.31

1CT (Average) 106.52 41.29 4.52 1910 884 203 40.5 117 680 15.3 25.4 0.340 1.30 74.7 58.3 45.3 51.3 51.0 54.0 37.0 47.6 55.2 47.4

Calculated Heads 257.99 100.00 1054 626 185 33 94 520 17.1 22.0 0.254 1.13 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 Accountabilities 88.9% 98.8% 121% 90.5% 73.7% 98.3% 103.3% 106.5% 96.6% 98.4%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm opt % % % % 1CC1 19.77 7.66 21.97 203 181 88 11 48 212 9.2 11.2 0.093 0.574 1.5 2.2 3.6 2.6 3.9 3.1 4.1 3.9 2.8 3.9 1CC1 - 1CC2 51.38 19.92 21.50 220 213 116 13 51 266.1 12.6 14.0 0.111 0.717 4.2 6.8 12.4 8.2 10.8 10.2 14.6 12.7 8.7 12.7 1CC1 - 1CC3 90.84 35.21 21.73 267 284 143 18 62 331 16.0 16.9 0.142 0.87 9.0 16.0 27.2 19.5 23.1 22.4 32.9 27.2 19.6 27.0 1CC1 - 1CC4 130.72 50.67 19.54 368 380 163 24 73 381 17.9 18.9 0.180 0.97 17.8 30.8 44.7 37.1 39.2 37.1 53.0 43.6 35.8 43.2 1CC1 - 1CC5 151.47 58.71 15.94 452 444 172 27 79 407 18.4 19.6 0.194 1.02 25.3 41.7 54.7 48.7 49.0 46.0 63.0 52.4 44.8 52.6

95

Table D-3: Flotation Data Sheet Z10-6-3

Composite: Plant Final Concentrate Date: 23-Jan-12 Z10 6-3 Plant Sample from Feb 10, 2009 Purpose: Baseline Test Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Penalty element separation baseline test series Regrind: RMA % 15.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell pH W22C H55 Stage(s) Grind Cond. Float mv wt (g) liters Regrind 15.0 7.97 194 Cond 5.0 7.98 -61 1CC1 17 34 1.0 94 2.3 1CC2 34 34 2.0 8.03 31 148 2.3 1CC3 51 3.0 8.04 60 193 2.3 1CC4 34 5.0 8.04 101 264 2.3 1CC5 5.0 8.06 112 213 2.3

CT= 2520

Total ###### 68.000 15.0 5.0 16.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm opt % % % % 1CC1 15.11 6.37 16.07 211 83 41 10 19 105 3.33 6.42 0.055 0.228 1.4 0.9 2.0 2.1 1.4 1.3 1.3 1.9 1.5 1.4 1CC2 23.14 9.76 15.64 211 203 87 14 44 210 9.25 12.0 0.091 0.546 2.2 3.4 6.6 4.5 5.0 4.1 5.4 5.4 3.7 5.0 1CC3 28.29 11.93 14.66 242 283 133 17 56 320 15.2 17.0 0.135 0.80 3.0 5.7 12.3 6.7 7.8 7.7 10.9 9.3 6.8 9.0 1CC4 33.41 14.09 12.66 290 436 178 25 79 428 20.8 22.1 0.211 1.04 4.3 10.5 19.4 11.7 12.9 12.1 17.7 14.3 12.5 13.8 1CC5 17.23 7.27 8.09 449 659 203 37 102 526 22.0 23.9 0.275 1.17 3.4 8.1 11.4 8.9 8.6 7.7 9.7 8.0 8.4 8.0

1CTA 1610 830 224 37 110 661 17.9 26.3 0.315 1.33 1CTB 1610 22 42 109 659 18.2 26.3 0.314 1.32

1CT (Average) 119.86 50.57 4.76 1610 830 123 39.5 110 660 18.1 26.3 0.314 1.33 85.7 71.4 48.3 66.1 64.3 67.1 55.0 61.1 67.1 62.8

Calculated Heads 237.04 100.00 951 588 129 30 86 498 16.6 21.8 0.237 1.07 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 Accountabilities 80.1% 92.9% 84% 83.9% 67.3% 94.1% 100.3% 105.4% 90.1% 92.6%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm opt % % % % 1CC1 15.11 6.37 16.07 211 83 41 10 19 105 3.3 6.4 0.055 0.228 1.4 0.9 2.0 2.1 1.4 1.3 1.3 1.9 1.5 1.9 1CC1 - 1CC2 38.25 16.14 15.81 211 156 69 12 34 168.5 6.9 9.8 0.076 0.420 3.6 4.3 8.6 6.6 6.4 5.4 6.7 7.3 5.2 7.3 1CC1 - 1CC3 66.54 28.07 15.30 224 210 96 14 43 233 10.4 12.9 0.101 0.58 6.6 10.0 20.9 13.3 14.2 13.1 17.6 16.6 12.0 15.4 1CC1 - 1CC4 99.95 42.17 14.30 246 285 123 18 55 298 13.9 15.9 0.138 0.74 10.9 20.5 40.3 25.0 27.1 25.2 35.3 30.9 24.5 29.2 1CC1 - 1CC5 117.18 49.43 12.85 276 340 135 21 62 332 15.1 17.1 0.158 0.80 14.3 28.6 51.7 33.9 35.7 32.9 45.0 38.9 32.9 37.2

96

Table D-4: Flotation Data Sheet Z10-8-1

Composite: Plant Final Concentrate Date: 23-Jan-12 Z10 8-1 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 5.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell pH Na S SIPX W22C H55 Stage(s) 2 Grind Cond. Float mv wt (g) liters Regrind 5.0 7.96 120 Cond ###### 5.0 11.93 -501 1CC1 34 34 1.0 11.94 -490 290 2.3 1CC2 17 34 2.0 11.98 -349 149 2.3 1CC3 34 3.0 11.99 -406 262 2.3 1CC4 ###### 34 5.0 11.99 -193 254 2.3 1CC5 51 5.0 12.05 -349 149 2.3 Scav 22.0 5.0 12.0 -323 131 2.3 CT= 2254

Total ###### ###### 68.000 5.0 5.0 21.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 25.76 10.70 8.88 109 98 11 8 13 44 1.3 4.3 0.048 0.1515 1.3 1.8 0.7 2.5 1.9 0.9 0.9 2.4 2.2 1.6 1CC2 20.01 8.31 13.43 105 94 22 7 11 50 1.2 4.4 0.043 0.1374 1.0 1.3 1.0 1.7 1.2 0.8 0.7 1.9 1.5 1.1 1CC3 11.09 4.60 4.23 293 169 40 11 25 97 2.1 10.3 0.067 0.22 1.5 1.3 1.0 1.5 1.6 0.9 0.6 2.4 1.3 1.0 1CC4 29.72 12.34 11.70 1516 588 210 35 93 546 13.7 26.1 0.334 1.08 20.6 12.5 14.6 12.9 15.5 13.4 10.8 16.5 17.8 12.8 1CC5 4.26 1.77 2.86 287 300 69 15 28 201 5.3 9.3 0.116 0.46 0.6 0.9 0.7 0.8 0.7 0.7 0.6 0.8 0.9 0.8 SVC 4.99 2.07 3.81 1363 45 129 28 75 354 7.0 26.7 0.16 0.45 3.1 0.2 1.5 1.7 2.1 1.5 0.9 2.8 1.4 0.9

1CTA 1078 791 239 46 94 684 22.2 23.8 0.289 1.40 1CTB 1097 785 236 42 96 677 22.1 23.7 0.288 1.44

1CT (Average) 145.01 60.21 6.43 1088 788.0 238 44.0 95 681 22.1 23.7 0.288 1.4 71.9 82.0 80.5 78.9 77.0 81.8 85.5 73.2 74.9 81.8

Calculated Heads 240.84 100.00 909 579 178 34 74 501 16 19.5 0.232 1 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1 Accountabilities 76.6% 91.5% 116% 93.3% 58.0% 94.8% 94.2% 94.7% 88.2% 90.8%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 25.76 10.70 8.88 109 98 11 8 13 44 1.3 4.3 0.048 0.152 1.3 1.8 0.7 2.5 1.9 0.9 0.9 2.4 2.2 2.4 1CC1 - 1CC2 45.77 19.00 10.43 107 96 16 8 12 46.6 1.2 4.4 0.046 0.145 2.3 3.1 1.7 4.2 3.1 1.7 1.6 4.3 3.7 4.3 1CC1 - 1CC3 56.86 23.61 8.11 143 110 21 8 15 56 1.4 5.5 0.050 0.160 3.8 4.4 2.7 5.7 4.7 2.6 2.2 6.7 5.0 3.7 1CC1 - 1CC4 86.58 35.95 9.07 615 274 86 17 42 224 5.6 12.6 0.147 0.477 24.4 16.9 17.3 18.6 20.2 16.0 13.0 23.2 22.8 16.5 1CC1 - 1CC5 90.84 37.72 8.23 599 276 85 17 41 223 5.6 12.4 0.146 0.477 25.0 17.8 18.0 19.4 20.9 16.7 13.6 24.0 23.7 17.3 1CC + SVC 95.83 39.79 7.76 639 264 87 18 43 230 6 13 0.147 0.475 28.1 18.0 19.5 21.1 23.0 18.2 14.5 26.8 25.1 18.2

97

Table D-5: Flotation Data Sheet Z10-8-2

Composite: Plant Final Concentrate Date: 23-Jan-12 Z10 8-2 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell pH Na S Act. C W22C H55 Stage(s) 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 8.21 111 Cond ###### 50.0 7.91 103 Cond ###### 5.0 11.64 -506 1CC1 34 1.0 11.76 -407 116 2.3 1CC2 34 2.0 11.81 -365 123 2.3 1CC3 1907.7 17 3.0 11.85 -354 162 2.3 1CC4 5511.3 34 5.0 11.91 -386 172 2.3 1CC5 6189.6 34 5.0 11.94 -367 139 2.3 CT= 2341

Total ###### 1271.8 ###### 0.000 0.0 55.0 16.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 28.22 11.96 24.33 108 99 15 7 13 32 0.7 4.3 0.055 0.1329 1.5 2.1 1.0 2.6 2.3 0.8 0.5 2.4 2.5 1.3 1CC2 19.88 8.43 16.16 107 108 17 6 12 39 0.9 4.8 0.053 0.1417 1.0 1.6 0.8 1.6 1.5 0.7 0.4 1.9 1.7 1.0 1CC3 9.97 4.23 6.15 180 178 34 11 21 78 1.7 7.9 0.078 0.20 0.9 1.3 0.8 1.5 1.3 0.7 0.4 1.5 1.2 0.7 1CC4 8.86 3.76 5.15 247 279 58 18 31 149 3.1 10.4 0.119 0.32 1.1 1.9 1.2 2.1 1.7 1.1 0.7 1.8 1.7 1.0 1CC5 3.68 1.56 2.65 475 410 108 29 58 237 4.9 20.7 0.164 0.43 0.8 1.1 0.9 1.4 1.3 0.8 0.4 1.5 1.0 0.5

1CTA 1185 743 254 40 89 687 24.7 28.7 0.350 1.64 1CTB 1201 730 241 43 90 660 24.0 28.0 0.344 1.71

1CT (Average) 165.27 70.07 7.06 1193 736.5 248 41.5 90 674 24.3 28.3 0.347 1.7 94.7 92.0 95.3 90.8 91.9 95.9 97.6 90.9 91.9 95.5

Calculated Heads 235.88 100.00 882 561 182 32 68 492 17 21.8 0.264 1 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1 Accountabilities 74.4% 88.7% 119% 88.9% 53.3% 92.9% 105.5% 105.8% 100.4% 106.9%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 28.22 11.96 24.33 108 99 15 7 13 32 0.7 4.3 0.055 0.133 1.5 2.1 1.0 2.6 2.3 0.8 0.5 2.4 2.5 2.4 1CC1 - 1CC2 48.10 20.39 20.13 108 103 16 7 13 34.9 0.8 4.5 0.054 0.137 2.5 3.7 1.8 4.2 3.8 1.5 0.9 4.3 4.2 4.3 1CC1 - 1CC3 58.07 24.62 14.48 120 116 19 7 14 42 0.9 5.1 0.058 0.148 3.4 5.0 2.6 5.7 5.1 2.2 1.3 5.8 5.4 3.0 1CC1 - 1CC4 66.93 28.37 11.68 137 137 24 9 16 56 1.2 5.8 0.066 0.171 4.5 6.9 3.8 7.8 6.8 3.3 2.0 7.6 7.1 4.0 1CC1 - 1CC5 70.61 29.93 9.92 154 151 28 10 18 66 1.4 6.6 0.071 0.184 5.3 8.0 4.7 9.2 8.1 4.1 2.4 9.1 8.1 4.5

98

Table D-6: Flotation Data Sheet Z10-8-3

Composite: Plant Final Concentrate Date: 23-Jan-12 Z10 8-3 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec pH Na S Act. C W22C H55 Stage(s) 2 404 Grind Cond. Float mv wt (g) liters Regrind 0.0 Cond ###### 50.0 7.87 82 Cond ###### 5.0 11.61 -456 1CC1 1067.4 22.0 34 1.0 11.86 -453 96 2.3 1CC2 22.0 34 2.0 11.89 -406 139 2.3 1CC3 2811.3 34 3.0 11.93 -387 171 2.3 1CC4 1720.4 22.0 34 5.0 11.94 -355 173 2.3 1CC5 7175.2 11.0 5.0 11.97 -357 217 2.3 CT= 2273

Total ###### 1258.8 ###### 0.000 0.0 55.0 16.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 20.57 8.63 21.43 101 93 6 6 12 22 0.7 3.9 0.048 0 0.9 1.4 0.3 1.6 1.4 0.4 0.3 1.5 1.8 0.9 1CC2 23.43 9.83 16.86 95 92 10 6 10 26 0.7 4.1 0.043 0 1.0 1.6 0.6 1.8 1.4 0.5 0.4 1.8 1.9 1.0 1CC3 12.15 5.10 7.11 145 155 25 8 18 67 1.6 6.3 0.067 0.19 0.8 1.4 0.7 1.3 1.3 0.7 0.5 1.5 1.5 0.8 1CC4 7.48 3.14 4.32 291 267 57 19 34 145 2.9 11.7 0.114 0.28 1.0 1.5 1.0 1.8 1.5 1.0 0.5 1.7 1.6 0.8 1CC5 6.06 2.54 2.79 441 410 100 24 51 240 5.0 17.7 0.162 0.45 1.2 1.8 1.5 1.9 1.8 1.3 0.7 2.0 1.8 1.0

1CTA 1364 706 231 39 94 636 24.7 28.9 0.249 1.60 1CTB 1176 788 238 45 93 649 24.4 27.9 0.343 1.55

1CT (Average) 168.63 70.76 7.42 1270 747.0 235 42.0 94 643 24.5 28.4 0.296 1.6 95.1 92.3 95.9 91.6 92.6 96.1 97.6 91.5 91.4 95.5

Calculated Heads 238.32 100.00 944 572 173 32 71 473 18 22.0 0.229 1 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1 Accountabilities 79.6% 90.4% 113% 90.1% 55.8% 89.4% 107.4% 106.5% 87.1% 101.5%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 20.57 8.63 21.43 101 93 6 6 12 22 0.7 3.9 0.048 0.118 0.9 1.4 0.3 1.6 1.4 0.4 0.3 1.5 1.8 1.5 1CC1 - 1CC2 44.00 18.46 18.72 98 92 8 6 11 24.1 0.7 4.0 0.045 0.121 1.9 3.0 0.9 3.4 2.8 0.9 0.7 3.3 3.7 3.3 1CC1 - 1CC3 56.15 23.56 13.83 108 106 12 6 12 33 0.9 4.5 0.050 0.136 2.7 4.4 1.6 4.7 4.1 1.6 1.2 4.8 5.2 2.7 1CC1 - 1CC4 63.63 26.70 10.99 130 125 17 8 15 47 1.1 5.3 0.058 0.153 3.7 5.9 2.6 6.5 5.6 2.6 1.7 6.5 6.8 3.5 1CC1 - 1CC5 69.69 29.24 8.76 157 150 24 9 18 63 1.5 6.4 0.067 0.179 4.9 7.7 4.1 8.4 7.4 3.9 2.4 8.5 8.6 4.5

99

Table D-7:Flotation Data Sheet Z10-8-4

Composite: Plant Final Concentrate Date: 9-Mar-12 Z10 8-4 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec pH Lime 7262 W22C H55 Stage(s) 3894 Grind Cond. Float mv wt (g) liters Regrind 0.0 6.92 200 Cond 45.0 11.18 -14 Cond 5202.9

1CC1 2 µl 40 5.0 10.83 33 2.3 1CC2 173.4 80 5.0 10.85 24 2.3

1CC3 173.4 1 µl 20 5.0 5.0 11.00 10 2.3 1CC4 520.3 5.0 10.85 24 2.3

Total 6070.1 140.0 0.000 0.000 0.0 50.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 82.49 35.77 232 328 147 37 27 286 14.2 16.2 0.118 1 8.3 18.8 32.2 25.6 16.3 20.9 32.0 26.8 17.5 32.4 1CC2 11.37 4.93 307 354 113 31 23 253 11.2 14.9 0.134 1 1.5 2.8 3.4 3.0 1.9 2.5 3.5 3.4 2.8 3.5 1CC3 16.91 7.33 578 479 200 38 48 462 19.9 23.8 0.190 1.22 4.2 5.6 9.0 5.4 5.9 6.9 9.2 8.1 5.8 8.3 1CC4 5.49 2.38 969 610 183 42 54 441 16.6 21.8 0.246 1.13 2.3 2.3 2.7 1.9 2.2 2.1 2.5 2.4 2.4 2.5

1CTA 1688 874 172 68 89 641 16.6 25.3 0.345 1.18 1CTB 1679 897 175 66 87 690 17.0 26.3 0.347 1.12

1CT (Average) 114.38 49.59 1684 885.5 174 67.0 88 666 16.8 25.8 0.346 1.1 83.7 70.5 52.7 64.1 73.7 67.6 52.8 59.3 71.5 53.3

Calculated Heads 230.64 100.00 998 624 163 52 59 489 16 21.6 0.240 1 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1 Accountabilities 84.2% 98.5% 107% 143.8% 46.3% 92.5% 95.6% 104.7% 91.3% 93.2%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 82.49 35.77 232 328 147 37 27 286 14.2 16.2 0.118 0.972 8.3 18.8 32.2 25.6 16.3 20.9 32.0 26.8 17.5 26.8 1CC1 - 1CC2 93.86 40.70 241 331 143 36 27 282.0 13.8 16.0 0.120 0.948 9.8 21.6 35.6 28.6 18.2 23.4 35.5 30.2 20.3 30.2 1CC1 - 1CC3 110.77 48.03 293 354 152 37 30 309 14.7 17.2 0.130 0.989 14.0 27.2 44.6 34.0 24.1 30.3 44.7 38.3 26.1 44.2 1CC1 - 1CC4 116.26 50.41 324 366 153 37 31 316 14.8 17.4 0.136 0.996 16.3 29.5 47.3 35.9 26.3 32.4 47.2 40.7 28.5 46.7

100

Table D-8: Flotation Data Sheet Z10-8-5

Composite: Plant Final Concentrate Date: 9-Mar-12 Z10 8-5 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime SO W22C H55 Stage(s) 3894 7262 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 7.1 61 Cond 2 µl 16740 Conditon with SO2 30.0 5.2 94

Cond ###### 2 µl 40 Condition with Air 30.0 11.0 -33 1CC1 5.0 10.7 -32 2.3 1CC2 544.0 40 5.0 10.9 -63 2.3 1CC3 5.0 5.0 10.6 -42 2.3 1CC4 5.0 10.3 -27 2.3

Total ###### 80.0 0.000 0.000 0.0 65.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 74.75 31.28 377 197 159 31 31 316 17.2 21.1 0.079 1 12.1 10.0 30.3 20.1 16.8 20.0 34.8 30.4 10.3 27.0 1CC2 5.96 2.49 489 267 108 19 13 187 6.3 10.1 0.119 1 1.3 1.1 1.6 1.0 0.6 0.9 1.0 1.2 1.2 1.7 1CC3 5.39 2.26 575 315 89 32 11 183 4.9 8.6 0.134 0.62 1.3 1.1 1.2 1.5 0.4 0.8 0.7 0.9 1.3 1.3 1CC4 4.16 1.74 644 303 78 30 9 168 4.4 8.2 0.131 0.52 1.2 0.9 0.8 1.1 0.3 0.6 0.5 0.7 0.9 0.9

1CTA 1303 868 174 60 78 617 15.7 23.5 0.334 1.17 1CTB 1322 860 174 58 74 615 15.6 23.1 0.332 1.17

1CT (Average) 148.69 62.23 1313 864.0 174 59.0 76 616 15.7 23.3 0.333 1.2 84.1 86.9 66.1 76.3 81.9 77.7 63.0 66.8 86.3 69.1

Calculated Heads 238.95 100.00 971 618 164 48 58 494 15 21.7 0.240 1 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1 Accountabilities 81.9% 97.7% 107% 133.7% 45.1% 93.4% 93.5% 105.1% 91.3% 91.7%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 74.75 31.28 377 197 159 31 31 316 17.2 21.1 0.079 0.910 12.1 10.0 30.3 20.1 16.8 20.0 34.8 30.4 10.3 30.4 1CC1 - 1CC2 80.71 33.78 385 202 155 30 30 306.4 16.4 20.2 0.082 0.895 13.4 11.1 31.9 21.1 17.4 20.9 35.8 31.6 11.5 31.6 1CC1 - 1CC3 86.10 36.03 397 209 151 30 29 299 15.7 19.5 0.085 0.878 14.7 12.2 33.1 22.6 17.8 21.7 36.5 32.5 12.8 30.0 1CC1 - 1CC4 90.26 37.77 409 214 148 30 28 293 15.1 19.0 0.087 0.862 15.9 13.1 33.9 23.7 18.1 22.3 37.0 33.2 13.7 30.9

101

Table D-9: Flotation Data Sheet Z10-8-6

Composite: Plant Final Concentrate Date: 9-Mar-12 Z10 8-6 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime W22C H55 Stage(s) 3894 7261A Grind Cond. Float mv wt (g) liters Regrind 0.0 6.9 114

Cond 2554.8 2 µl 30.0 11.0 -38 Cond 80 10.0 11.0 5 1CC1 5.0 10.7 -7 2.3 1CC2 562.9 40 5.0 10.9 -37 2.3

1CC3 173.2 1 µl 5.0 10.8 31 2.3 1CC4 173.2 5.0 10.7 -23 2.3

Total 3464.1 120.0 0.000 0.000 0.0 40.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 87.82 38.03 275 308 155 22 32 342 15.1 18.3 0.145 0.918 10.6 19.3 35.6 17.1 20.5 26.3 33.8 31.6 23.5 34.1 1CC2 8.81 3.81 257 378 124 35 25 274 9.3 12.6 0.171 0.772 1.0 2.4 2.9 2.7 1.6 2.1 2.1 2.2 2.8 2.9 1CC3 8.63 3.74 441 494 220 49 50 482 17.5 20.2 0.236 1.31 1.7 3.0 5.0 3.7 3.1 3.6 3.8 3.4 3.8 4.8 1CC4 7.96 3.45 507 579 217 42 49 510 16.4 20.3 0.263 1.30 1.8 3.3 4.5 3.0 2.8 3.6 3.3 3.2 3.9 4.4

1CTA 1648 858 166 74 85 609 18.4 25.3 0.305 1.08 1CTB 1648 172 67 83 637 19.7 26.1 0.301 1.08

1CT (Average) 117.72 50.97 1648 858.0 169 70.5 84 623 19.0 25.7 0.303 1.1 84.9 72.0 52.0 73.5 72.0 64.4 57.0 59.6 66.0 53.8

Calculated Heads 230.94 100.00 988 607 166 49 59 493 17 22.0 0.234 1 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1 Accountabilities 83.3% 95.9% 108% 135.9% 46.5% 93.3% 102.8% 106.5% 89.0% 89.0%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 87.82 38.03 275 308 155 22 32 342 15.1 18.3 0.145 0.918 10.6 19.3 35.6 17.1 20.5 26.3 33.8 31.6 23.5 31.6 1CC1 - 1CC2 96.63 41.84 273 314 152 23 31 335.7 14.6 17.7 0.147 0.905 11.6 21.7 38.5 19.8 22.1 28.4 35.9 33.8 26.3 33.8 1CC1 - 1CC3 105.26 45.58 287 329 158 25 33 348 14.8 17.9 0.154 0.938 13.3 24.7 43.5 23.5 25.2 32.0 39.7 37.2 30.1 41.8 1CC1 - 1CC4 113.22 49.03 303 347 162 26 34 359 14.9 18.1 0.162 0.963 15.1 28.0 48.0 26.5 28.0 35.6 43.0 40.4 34.0 46.2

102

Table D-10: Flotation Data Sheet Z10-8-7

Composite: Plant Final Concentrate Date: 9-Mar-12 Z10 8-7 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime SO W22C H55 Stage(s) 3894 7261A 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 6.9 133

Cond 2 µl 47934 30.0 2.1 192 Cond ###### 80 30.0 11.3 -285 1CC1 5.0 11.0 -44 2.3 1CC2 40 5.0 11.0 -33 2.3

1CC3 1 µl 5.0 10.6 -12 2.3 1CC4 5.0 10.4 -5 2.3

Total ###### 120.0 0.000 0.000 0.0 60.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 19.90 8.38 232 240 54 27 11 151 6.2 9.9 0.123 0.366 2.0 3.4 2.9 4.1 1.7 2.6 2.8 3.8 5.1 3.0 1CC2 18.08 7.62 340 222 132 25 33 328 18.9 19.8 0.086 0.803 2.7 2.9 6.5 3.5 4.6 5.1 7.8 6.8 3.2 5.9 1CC3 9.08 3.82 317 231 109 20 14 207 10.6 12.6 0.096 0.64 1.3 1.5 2.7 1.4 1.0 1.6 2.2 2.2 1.8 2.4 1CC4 5.98 2.52 274 246 87 20 10 174 7.2 9.8 0.105 0.55 0.7 1.1 1.4 0.9 0.5 0.9 1.0 1.1 1.3 1.3

1CTA 1137 696 172 65 67 566 20.6 24.6 0.231 1.16 1CTB 1168 686 171 63 62 570 20.4 24.5 0.233 1.16

1CT (Average) 184.37 77.66 1153 691.0 172 64.0 65 568 20.5 24.5 0.232 1.2 93.3 91.1 86.5 90.1 92.2 89.8 86.2 86.1 88.6 87.4

Calculated Heads 237.41 100.00 959 589 154 55 54 491 18 22.1 0.203 1 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1 Accountabilities 80.9% 93.0% 101% 153.2% 42.4% 92.8% 111.6% 107.2% 77.2% 89.7%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 19.90 8.38 232 240 54 27 11 151 6.2 9.9 0.123 0.366 2.0 3.4 2.9 4.1 1.7 2.6 2.8 3.8 5.1 3.8 1CC1 - 1CC2 37.98 16.00 283 231 91 26 21 235.2 12.2 14.6 0.105 0.574 4.7 6.3 9.4 7.6 6.3 7.7 10.6 10.6 8.3 10.6 1CC1 - 1CC3 47.06 19.82 290 231 95 25 20 230 11.9 14.2 0.104 0.587 6.0 7.8 12.1 9.0 7.3 9.3 12.8 12.8 10.1 11.3 1CC1 - 1CC4 53.04 22.34 288 233 94 24 19 224 11.4 13.7 0.104 0.582 6.7 8.9 13.5 9.9 7.8 10.2 13.8 13.9 11.4 12.6

103

Table D-11: Flotation Data Sheet Z10-8-8

Composite: Plant Rougher Concentrate Date: 9-Mar-12 Z10 8-8 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 270 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec TX pH Lime W22C H55 Stage(s) 3894 15155 Grind Cond. Float mv wt (g) liters Regrind 0.0 6.4 183

Cond 3401.4 2 µl 30.0 11.0 -20

Cond 2 µl 5.0 11.0 -7 1CC1 5.0 10.6 -8 2.3

1CC2 591.5 2 µl 5.0 10.9 -3 2.3

1CC3 295.8 2 µl 2 µl 5.0 10.9 -2 2.3 1CC4 5.0 10.7 -22 2.3

Total 4288.7 0.0 0.000 0.000 0.0 35.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 95.05 35.14 4985 1175 602 82 96 703 27.7 32.4 0.620 4 10.7 47.3 68.6 31.3 35.5 52.3 70.3 35.5 56.5 54.3 1CC2 28.97 10.71 17295 1241 335 148 138 722 17.0 34.5 0.483 4 11.3 15.2 11.6 17.2 15.6 16.4 13.2 11.6 13.4 15.6 1CC3 15.11 5.59 23644 1118 248 133 134 668 15.0 36.1 0.402 4.12 8.1 7.2 4.5 8.1 7.9 7.9 6.1 6.3 5.8 8.1 1CC4 6.31 2.33 30890 1137 226 141 138 648 11.9 34.7 0.408 4.05 4.4 3.0 1.7 3.6 3.4 3.2 2.0 2.5 2.5 3.3

1CTA 23546 517 91 82 72 209 2.58 30.7 0.185 1.15 1CTB 23000 512 90 77 83 205 2.48 30.4 0.182 1.14 0.0 1CT (Average) 125.04 46.23 23273 514.5 91 79.5 78 207 2.5 30.6 0.183 1.1 65.5 27.3 13.6 39.8 37.6 20.2 8.4 44.1 21.8 18.7

Calculated Heads 270.48 100.00 16405 873 308 92 95 473 13.8 32.0 0.386 3 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3 Accountabilities 87.1% 96.6% 101% 121.2% 60.9% 88.6% 97.7% 104.7% 85.8% 91.5%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 95.05 35.14 4985 1175 602 82 96 703 27.7 32.4 0.620 4.380 10.7 47.3 68.6 31.3 35.5 52.3 70.3 35.5 56.5 35.5 1CC1 - 1CC2 124.02 45.85 7861 1190 540 97 106 707.4 25.2 32.9 0.588 4.324 22.0 62.5 80.2 48.5 51.1 68.7 83.5 47.1 69.9 47.1 1CC1 - 1CC3 139.13 51.44 9575 1183 508 101 109 703 24.1 33.2 0.568 4.302 30.1 69.7 84.7 56.6 59.0 76.6 89.6 53.4 75.7 78.0 1CC1 - 1CC4 145.44 53.77 10499 1181 496 103 110 701 23.5 33.3 0.561 4.291 34.5 72.7 86.4 60.2 62.4 79.8 91.6 55.9 78.2 81.3

104

Table D-12: Flotation Data Sheet Z10-8-9

Composite: Plant Rougher Concentrate Date: 9-Mar-12 Z10 8-9 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 270 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec TX pH Lime W22C H55 Stage(s) 3894 15281 Grind Cond. Float mv wt (g) liters Regrind 0.0 6.6 183 Cond 5687.1 30.0 11.0 19

Cond 2 µl 2 µl 5.0 11.0 -9 1CC1 5.0 10.7 3 2.3

1CC2 480.1 2 µl 5.0 10.7 1 2.3

1CC3 295.4 2 µl 2 µl 5.0 10.8 5 2.3 1CC4 5.0 10.5 18 2.3

Total 6462.6 0.0 0.000 0.000 0.0 35.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 98.41 36.34 5614 1141 598 107 135 707 28.2 32.4 0.605 4 12.3 47.4 69.7 39.6 45.6 53.8 72.4 60.4 56.4 55.7 1CC2 26.24 9.69 17917 1221 334 130 130 695 17.9 35.8 0.460 4 10.5 13.5 10.4 12.8 11.7 14.1 12.2 17.8 11.4 13.6 1CC3 15.10 5.58 22417 1051 237 134 125 612 14.1 35.5 0.360 3.69 7.6 6.7 4.2 7.6 6.5 7.1 5.6 10.1 5.1 7.2 1CC4 8.21 3.03 31234 1222 241 141 143 659 10.9 34.1 0.445 4.27 5.7 4.2 2.3 4.4 4.0 4.2 2.3 5.3 3.5 4.6

1CTA 23383 537 92 80 79 213 2.37 2.9 0.203 1.21 1CTB 23218 549 91 74 74 227 2.29 2.6 0.202 1.16

1CT (Average) 122.83 45.36 23301 543 92 77.0 77 220 2.3 2.8 0.202 1 63.9 28.2 13.4 35.6 32.2 20.8 7.5 6.4 23.6 18.9

Calculated Heads 270.79 100.00 16543 875 312 98 108 478 14.2 19.5 0.390 3 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3 Accountabilities 87.8% 96.9% 103% 129.2% 69.0% 89.7% 100.1% 63.7% 86.7% 91.7%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 98.41 36.34 5614 1141 598 107 135 707 28.2 32.4 0.605 4.360 12.3 47.4 69.7 39.6 45.6 53.8 72.4 60.4 56.4 60.4 1CC1 - 1CC2 124.65 46.03 8204 1158 542 112 134 704.6 26.0 33.1 0.575 4.282 22.8 60.9 80.1 52.4 57.3 67.9 84.6 78.2 67.8 78.2 1CC1 - 1CC3 139.75 51.61 9740 1146 509 114 133 695 24.8 33.4 0.551 4.218 30.4 67.6 84.3 60.0 63.8 75.0 90.2 88.3 72.9 76.5 1CC1 - 1CC4 147.96 54.64 10932 1150 495 116 134 693 24.0 33.4 0.545 4.221 36.1 71.8 86.6 64.4 67.8 79.2 92.5 93.6 76.4 81.1

105

Table D-13: Flotation Data Sheet Z10-8-10

Composite: Plant Rougher Concentrate Date: 9-Mar-12 Z10 8-10 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 270 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec TX pH Lime SO W22C H55 Stage(s) 3894 15281 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 6.4 157 Cond 17435 Pre-aeration w/ SO2 30.0 2.3 215 Cond ###### Pre-aeration w/ air 30.0 11.0 -505

1CC1 2 µl 2 µl 5.0 10.8 29 2.3 1CC2 484.3 5.0 10.6 14 2.3

1CC3 380.5 2 µl 5.0 10.7 -18 2.3 1CC4 5.0 10.7 -18 2.3

Total ###### 0.0 0.000 0.000 0.0 60.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 60.32 20.87 4199 632 487 12 54 690 27.9 28.7 0.493 3 5.2 13.5 29.0 4.2 15.3 29.1 44.8 19.0 22.9 20.4 1CC2 19.34 6.69 6986 1191 878 53 79 775 23.5 30.4 0.713 6 2.8 8.2 16.7 6.0 7.2 10.5 12.1 6.5 10.7 12.6 1CC3 17.34 6.00 10553 2125 1147 120 140 879 19.0 30.4 1.052 7.05 3.7 13.1 19.6 12.1 11.4 10.7 8.8 5.8 14.1 13.9 1CC4 8.21 2.84 14190 1982 693 131 130 884 15.2 30.7 0.929 5.47 2.4 5.8 5.6 6.3 5.0 5.1 3.3 2.8 5.9 5.1

1CTA 22774 936 155 64 73 339 6.38 33.1 0.330 2.30 1CTB 23146 888 166 69 69 355 6.27 32.1 0.325 2.30

1CT (Average) 183.87 63.61 22960 912.3 161 66.5 71 347 6.3 32.6 0.327 2.3 85.9 59.4 29.1 71.4 61.1 44.6 31.0 65.9 46.4 48.0

Calculated Heads 289.08 100.00 16983 975 351 59 74 494 13 31.5 0.448 3 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3 Accountabilities 90.1% 108.0% 115.4% 78.0% 47.3% 92.7% 91.7% 102.8% 99.6% 98.2%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 60.32 20.87 4199 632 487 12 54 690 27.9 28.7 0.493 2.975 5.2 13.5 29.0 4.2 15.3 29.1 44.8 19.0 22.9 19.0 1CC1 - 1CC2 79.66 27.56 4876 767 582 22 60 710.4 26.8 29.1 0.546 3.644 8.0 21.7 45.7 10.2 22.5 39.6 56.9 25.5 33.6 25.5 1CC1 - 1CC3 97.00 33.55 5891 1010 683 39 74 740 25.4 29.4 0.637 4.252 11.7 34.8 65.3 22.3 33.9 50.3 65.7 31.3 47.7 46.9 1CC1 - 1CC4 105.21 36.39 6538 1086 684 47 79 752 24.6 29.5 0.659 4.347 14.1 40.6 70.9 28.6 38.9 55.4 69.0 34.1 53.6 52.0

106

Table D-14: Flotation Data Sheet Z10-8-11

Composite: Plant Rougher Concentrate Date: 9-Mar-12 Z10 8-11 Plant Sample from Feb 10, 2009 Purpose: Reagent screening Grind:~80% passing µm Sands ~80% passing µm Feed: 270 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec TX pH Lime SO W22C H55 Stage(s) 3894 15155 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 6.7 156 Cond 10783 Pre-aeration w/ SO2 30.0 2.8 215 Cond ###### Pre-aeration w/ air 30.0 11.0 -78

1CC1 2 µl 5.0 10.7 -14 2.3 1CC2 359.4 5.0 10.8 -17 2.3

1CC3 251.6 2 µl 5.0 10.8 -10 2.3 1CC4 467.3 5.0 10.9 17 2.3

Total ###### 0.0 0.000 0.000 0.0 60.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 38.41 13.81 2492 586 757 22 60 544 28.7 29.5 0.371 5 2.0 8.0 31.0 5.0 10.4 15.3 28.9 12.6 11.7 23.5 1CC2 12.35 4.44 4829 1180 630 54 96 675.1 25.9 28.7 0.649 5 1.3 5.2 8.3 3.9 5.3 6.1 8.4 3.9 6.6 6.3 1CC3 29.83 10.72 5523 1402 763 66 111 832.5 25.7 30.0 0.770 5.43 3.5 14.9 24.2 11.6 14.9 18.2 20.1 9.9 18.9 18.3 1CC4 26.39 9.49 12408 1708 424 98 135 887.5 21.9 33.1 0.924 4.27 6.9 16.0 11.9 15.3 16.1 17.2 15.2 9.7 20.0 12.7

1CTA 23828 849 134 59 62 342.6 6.04 33.6 0.305 2.01 1CTB 23707 989 136 68 76 342.8 6.12 33.7 0.306 2.03

1CT (Average) 171.24 61.55 23768 919.3 135.0 63.5 69 343 6.1 33.6 0.306 2.0 86.3 55.9 24.6 64.2 53.3 43.2 27.4 63.9 42.8 39.2

Calculated Heads 278.22 100.00 16956 1011 338 61 80 489 14 32.4 0.438 3 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3 Accountabilities 90.0% 112.0% 111% 80.1% 51.1% 91.8% 96.7% 105.9% 97.3% 102.5%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 38.41 13.81 2492 586 757 22 60 544 28.7 29.5 0.371 5.410 2.0 8.0 31.0 5.0 10.4 15.3 28.9 12.6 11.7 12.6 1CC1 - 1CC2 50.76 18.24 3061 731 726 30 69 575.9 28.0 29.3 0.438 5.192 3.3 13.2 39.3 8.9 15.7 21.4 37.3 16.5 18.3 16.5 1CC1 - 1CC3 80.59 28.97 3972 979 740 43 84 671 27.1 29.6 0.561 5.280 6.8 28.1 63.5 20.5 30.6 39.6 57.4 26.4 37.2 48.1 1CC1 - 1CC4 106.98 38.45 6053 1159 662 57 97 724 25.8 30.4 0.651 5.031 13.7 44.1 75.4 35.8 46.7 56.8 72.6 36.1 57.2 60.8

107

Table D-15: Flotation Data Sheet Z10-8-12

Composite: Plant Rougher Concentrate Date: 9-Mar-12 Z10 8-12 Plant Sample from Feb 10, 2009 Purpose: Baseline Grind:~80% passing µm Sands ~80% passing µm Feed: 270 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec TX pH Lime SO W22C H55 Stage(s) 3894 15155 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 Cond Cond 5.0 7.5 70 1CC1 5.0 7.7 97 2.3 1CC2 5.0 7.7 112 2.3 1CC3 5.0 7.7 118 2.3 1CC4 5.0 7.7 116 2.3

Total 0.0 0.0 0.000 0.000 0.0 5.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 89.88 43.06 13567 752 353 45 86 448.4 17.0 35.9 0.398 2.81 35.9 34.3 51.4 35.2 42.8 40.8 55.1 48.6 40.2 39.5 1CC2 26.67 12.78 18989 1110 367 58 108 495 13.5 37.4 0.533 3.16 14.9 15.0 15.8 13.5 15.9 13.4 13.0 15.1 16.0 13.2 1CC3 19.62 9.40 21249 1255 329 75 98 556.9 13.3 40.3 0.515 3.52 12.3 12.5 10.4 12.8 10.6 11.1 9.4 11.9 11.4 10.8 1CC4 13.73 6.58 21654 1304 314 78 95 540.8 12.2 35.4 0.463 3.75 8.8 9.1 7.0 9.3 7.2 7.5 6.0 7.3 7.2 8.1

1CTA 16023 988 162 59 72 459.6 7.88 19.4 0.382 3.08 1CTB 16378 957 161 55 72 452.0 7.55 19.0 0.386 3.11

1CT (Average) 58.81 28.18 16201 972.4 161.5 57.0 72 456 7.72 19.2 0.384 3.09 28.1 29.1 15.4 29.2 23.5 27.2 16.5 17.1 25.2 28.4

Calculated Heads 208.71 100.00 16256 943 296 55 87 473 13.3 31.7 0.426 3.06 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 86.3% 104.5% 97% 72.4% 55.5% 88.7% 93.7% 103.7% 94.7% 98.8%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 89.88 43.06 13567 752 353 45 86 448 17.0 35.9 0.398 2.809 35.9 34.3 51.4 35.2 42.8 40.8 55.1 48.6 40.2 48.6 1CC1 - 1CC2 116.55 55.84 14808 834 356 48 91 459.1 16.2 36.2 0.429 2.889 50.8 49.3 67.2 48.7 58.7 54.2 68.1 63.7 56.2 63.7 1CC1 - 1CC3 136.17 65.24 15736 895 352 52 92 473 15.8 36.8 0.441 2.980 63.1 61.8 77.6 61.5 69.3 65.3 77.5 75.6 67.6 63.5 1CC1 - 1CC4 149.90 71.82 16278 932 349 54 92 479 15.4 36.7 0.443 3.050 71.9 70.9 84.6 70.8 76.5 72.8 83.5 82.9 74.8 71.6

108

Table D-16: Flotation Data Sheet Z10-8-13

Composite: Plant Final Concentrate Date: 9-Mar-12 Z10 8-13 Plant Sample from Feb 10, 2009 Purpose: Baseline Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime SO W22C H55 Stage(s) 3894 7261A 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 Cond Cond 5.0 7.8 74 1CC1 5.0 8.0 91 2.3 1CC2 5.0 8.0 105 2.3 1CC3 5.0 8.0 105 2.3 1CC4 5.0 7.9 111 2.3

Total 0.0 0.0 0.000 0.000 0.0 5.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 58.34 32.64 183 233 56 7 21 194 7.54 9.74 0.107 0.374 6.3 11.5 12.0 9.1 11.5 12.5 14.0 14.2 14.6 11.7 1CC2 26.59 14.88 442 576 169 28 62 529 25.3 25.2 0.236 1.14 6.9 13.0 16.6 16.6 15.5 15.6 21.4 16.7 14.7 16.2 1CC3 13.26 7.42 670 797 235 39 64 632 26.2 26.7 0.282 1.59 5.2 9.0 11.5 11.5 8.0 9.3 11.1 8.9 8.8 11.3 1CC4 12.21 6.83 1038 916 261 41 66 675 28.3 30.1 0.316 1.67 7.4 9.5 11.8 11.1 7.6 9.1 11.0 9.2 9.0 10.9

1CTA 1886 983 192 38 98 698 18.9 29.7 0.329 1.35 1CTB 1825 977 190 30 81 719 20.1 30.1 0.330 1.38

1CT (Average) 68.32 38.23 1856 980.0 191 34.0 90 709 19.5 29.9 0.330 1.4 74.2 57.0 48.1 51.7 57.4 53.5 42.5 51.0 52.9 49.9

Calculated Heads 178.72 100.00 955 658 152 25 60 506 17.6 22.4 0.239 1.05 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 Accountabilities 80.6% 103.9% 99% 69.8% 46.5% 95.6% 106.1% 108.6% 90.9% 91.0%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 58.34 32.64 183 233 56 7 21 194 7.5 9.7 0.107 0.374 6.3 11.5 12.0 9.1 11.5 12.5 14.0 14.2 14.6 14.2 1CC1 - 1CC2 84.93 47.52 264 340 91 14 34 298.8 13.1 14.6 0.147 0.614 13.2 24.5 28.6 25.7 27.0 28.1 35.4 30.9 29.3 30.9 1CC1 - 1CC3 98.19 54.94 319 402 111 17 38 344 14.9 16.2 0.165 0.746 18.4 33.5 40.1 37.2 35.0 37.4 46.5 39.8 38.1 39.2 1CC1 - 1CC4 110.40 61.77 398 459 127 20 41 380 16.4 17.8 0.182 0.848 25.8 43.0 51.9 48.3 42.6 46.5 57.5 49.0 47.1 50.1

109

Table D-17: Flotation Data Sheet Z10-8-14

Composite: Plant Rougher Concentrate Date: 9-Mar-12 Z10 8-14 Plant Sample from Feb 10, 2009 Purpose: Reagent Screening Grind:~80% passing µm Sands ~80% passing µm Feed: 270 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime W22C H55 Stage(s) 7261A 3894 Grind Cond. Float mv wt (g) liters Regrind 0.0 7.7 42 Cond 2839 30.0 1.0 -52

Cond 100.0 2 µL 10.0 11.1 -14 1CC1 5.0 10.7 6 2.3

1CC2 722 100.0 2 µL 5.0 10.9 -3 2.3 1CC3 529 100.0 5.0 11.0 -4 2.3 1CC4 100.0 5.0 10.7 -15 2.3

Total 4090.3 0.0 0.000 0.000 0.0 40.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 42.09 20.25 4362 915 709 44 74 623.5 26.3 31.3 0.557 5.37 5.4 19.2 48.0 14.2 16.3 25.5 41.2 19.7 26.1 37.1 1CC2 23.78 11.44 8339 1315 476 72 109 832.7 24.9 30.0 0.727 4.47 5.8 15.6 18.2 13.1 13.5 19.2 22.1 10.6 19.2 17.5 1CC3 4.11 1.98 14519 1629 364 96 112 812.6 14.2 26.8 0.790 4.72 1.7 3.3 2.4 3.0 2.4 3.2 2.2 1.6 3.6 3.2 1CC4 2.97 1.43 17121 1613 325 88 107 750.9 12.5 27.7 0.706 4.58 1.5 2.4 1.6 2.0 1.7 2.2 1.4 1.2 2.3 2.2

1CTA 21626 906 141 64 87 383.3 6.58 33.3 0.330 1.83 1CTB 21832 857 135 67 101 377.8 6.66 33.1 0.321 1.78

1CT (Average) 134.86 64.90 21729 881.6 138.0 65.5 94 381 6.62 33.2 0.326 1.80 85.6 59.5 29.8 67.7 66.1 49.9 33.1 66.9 48.8 40.0

Calculated Heads 207.81 100.00 16471 963 299 63 92 495 12.9 32.2 0.433 2.93 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 87.4% 106.7% 99% 82.7% 59.1% 92.9% 91.4% 105.3% 96.2% 94.5%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 42.09 20.25 4362 915 709 44 74 624 26.3 31.3 0.557 5.369 5.4 19.2 48.0 14.2 16.3 25.5 41.2 19.7 26.1 19.7 1CC1 - 1CC2 65.87 31.70 5798 1059 625 54 87 699.0 25.8 30.8 0.618 5.046 11.2 34.8 66.2 27.3 29.8 44.7 63.3 30.3 45.3 30.3 1CC1 - 1CC3 69.98 33.67 6310 1093 610 57 88 706 25.1 30.6 0.629 5.027 12.9 38.1 68.6 30.3 32.2 47.9 65.5 31.9 48.9 57.8 1CC1 - 1CC4 72.95 35.10 6750 1114 598 58 89 708 24.6 30.5 0.632 5.008 14.4 40.5 70.2 32.3 33.9 50.1 66.9 33.1 51.2 60.0

110

Table D-18: Flotation Data Sheet Z10-8-15

Composite: Plant Rougher Concentrate Date: 9-Mar-12 Z10 8-15 Plant Sample from Feb 10, 2009 Purpose: Reagent Screening Grind:~80% passing µm Sands ~80% passing µm Feed: 270 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime W22C H55 Stage(s) 7262 3894 Grind Cond. Float mv wt (g) liters Regrind 0.0 7.7 68 Cond 3567 30.0 11.0 38

Cond 100.0 2 µL 10.0 11.1 47 1CC1 5.0 10.8 0 2.3

1CC2 563 100.0 2 µL 5.0 10.9 0 2.3 1CC3 563 100.0 5.0 10.9 -3 2.3 1CC4 5.0 10.6 10 2.3

Total 4693.5 0.0 0.000 0.000 0.0 40.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 33.39 15.67 3392 822 795 27 85 587.2 26.6 28.9 0.509 6.19 3.2 13.4 41.1 7.2 14.5 18.7 31.1 14.1 18.3 32.6 1CC2 17.65 8.28 6081 1123 509 46 92 811.4 27.4 29.9 0.656 4.39 3.0 9.7 13.9 6.5 8.3 13.7 16.9 7.7 12.5 12.2 1CC3 9.94 4.67 8405 1188 461 63 103 810.8 25.2 29.8 0.703 4.45 2.4 5.8 7.1 5.0 5.2 7.7 8.7 4.3 7.5 7.0 1CC4 8.41 3.95 11282 1382 373 74 119 753.6 21.8 31.1 0.733 4.18 2.7 5.7 4.9 5.0 5.1 6.1 6.4 3.8 6.6 5.5

1CTA 21568 921 147 64 87 391.2 6.73 33.2 0.355 1.85 1CTB 21910 942 150 68 95 393.8 7.97 33.7 0.358 1.91

1CT (Average) 143.67 67.43 21739 931.5 148.5 66.0 91 393 7.35 33.4 0.356 1.88 88.7 65.4 33.0 76.3 66.9 53.8 36.9 70.1 55.1 42.7

Calculated Heads 213.06 100.00 16532 960 303 58 92 491 13.4 32.2 0.436 2.98 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 87.7% 106.3% 100% 76.9% 58.9% 92.2% 95.0% 105.1% 96.9% 96.0%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 33.39 15.67 3392 822 795 27 85 587 26.6 28.9 0.509 6.195 3.2 13.4 41.1 7.2 14.5 18.7 31.1 14.1 18.3 14.1 1CC1 - 1CC2 51.04 23.96 4322 926 696 34 87 664.7 26.9 29.2 0.560 5.571 6.2 23.1 55.0 13.7 22.8 32.4 48.0 21.8 30.8 21.8 1CC1 - 1CC3 60.98 28.62 4987 969 658 38 90 689 26.6 29.3 0.583 5.389 8.6 28.9 62.1 18.7 28.0 40.1 56.7 26.1 38.3 51.8 1CC1 - 1CC4 69.39 32.57 5750 1019 623 43 93 696 26.0 29.5 0.601 5.243 11.3 34.6 67.0 23.7 33.1 46.2 63.1 29.9 44.9 57.3

111

Table D-19: Flotation Data Sheet Z10-8-16

Composite: Plant Final Concentrate Date: 9-Mar-12 Z10 8-16 Plant Sample from Feb 10, 2009 Purpose: Reagent Screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added, (g/t) Flotation Time (min) Pt Conc Cell TX Cytec pH Lime W22C H55 Stage(s) 15281 3894 Grind Cond. Float mv wt (g) liters Regrind 0.0 8.0 105 Cond 2926 30.0 11.1 -12

Cond 100 2µL 10.0 11.0 2 1CC1 5.0 10.7 17 2.3

1CC2 1069 100 2µL 5.0 10.9 3 2.3 1CC3 225 100 5.0 10.7 28 2.3 1CC4 563 5.0 11.0 7 2.3

Total 4782.3 0.0 0.000 0.000 0.0 40.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 136.07 76.56 827 595 161 27 55 509 20.0 21.9 0.208 1.26 64.8 69.1 85.4 76.8 78.3 77.7 88.3 77.9 73.5 86.0 1CC2 19.61 11.03 2134 1137 140 42 63 678 13.5 26.5 0.323 1.05 24.1 19.1 10.7 17.2 12.9 14.9 8.6 13.5 16.5 10.3 1CC3 6.76 3.80 1327 943 78 26 54 498 8.7 18.9 0.238 0.650 5.2 5.4 2.1 3.7 3.8 3.8 1.9 3.3 4.2 2.2 1CC4 1.47 0.83 932 708 48 20 34 334 5.7 14.8 0.195 0.438 0.8 0.9 0.3 0.6 0.5 0.6 0.3 0.6 0.7 0.3

1CTA 669 483 31 6 30 202 2.2 13.6 0.142 0.174 1CTB 624 440 27 6 31 191 1.8 12.5 0.133 0.146

1CT (Average) 13.83 7.78 647 461.5 29 6.0 31 196 2.0 13.0 0.137 0.160 5.1 5.5 1.5 1.7 4.5 3.0 0.9 4.7 5.1 1.2

Calculated Heads 177.74 100.00 977 658 144 27 54 502 17 21.6 0.216 1.12 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 Accountabilities 82.4% 104.0% 94% 74.8% 42.0% 94.8% 104.6% 104.5% 82.1% 97.5%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 136.07 76.56 827 595 161 27 55 509 20.0 21.9 0.208 1.260 64.8 69.1 85.4 76.8 78.3 77.7 88.3 77.9 73.5 77.9 1CC1 - 1CC2 155.68 87.59 992 663 158 29 56 530.6 19.2 22.5 0.222 1.233 88.9 88.2 96.1 94.0 91.2 92.6 96.9 91.4 90.0 91.4 1CC1 - 1CC3 162.44 91.39 1006 675 155 29 56 529 18.7 22.4 0.223 1.209 94.1 93.6 98.2 97.7 95.0 96.4 98.8 94.7 94.2 98.5 1CC1 - 1CC4 163.91 92.22 1005 675 154 29 56 527 18.6 22.3 0.222 1.202 94.9 94.5 98.5 98.3 95.5 97.0 99.1 95.3 94.9 98.8

112

Table D20: Flotation Data Sheet Z10-10-1

Composite: Plant Final Concentrate Date: 9-Mar-12 Z10 10-1 Plant Sample from Feb 10, 2009 Purpose: Reagent Screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added Flotation Time (min) Pt Conc Cell TX Cytec pH Lime W22C H55 Stage(s) 15155 3894 Grind Cond. Float mv wt (g) liters Regrind 0.0 8.1 157 Cond 3172 30.0 11.2 -2

Cond 100 2µL 10.0 11.1 -6 1CC1 5.0 10.8 8 2.3

1CC2 323 100 2µL 5.0 10.8 16 2.3 1CC3 161 100 5.0 11.1 3 2.3 1CC4 5.0 10.9 7 2.3

Total 3655.5 0.0 0.000 0.000 0.0 40.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 144.57 77.72 828 622 181 28 59 534 20.2 23.1 0.268 1.25 67.6 72.2 89.3 80.8 78.8 81.6 90.7 81.2 80.4 88.4 1CC2 20.30 10.91 1900 1059 112 34 76 564 10.7 21.9 0.297 0.862 21.8 17.3 7.8 13.8 14.3 12.1 6.7 10.8 12.5 8.6 1CC3 5.62 3.02 1448 896 68 21 51 467 8.2 21.0 0.219 0.562 4.6 4.0 1.3 2.4 2.6 2.8 1.4 2.9 2.6 1.5 1CC4 1.89 1.02 1243 746 62 8 41 350 5.6 17.7 0.187 0.518 1.3 1.1 0.4 0.3 0.7 0.7 0.3 0.8 0.7 0.5

1CTA 598 491 28 7 25 191 1.9 13.0 0.140 0.155 1CTB 631 493 26 13 31 195 1.9 12.8 0.129 0.153

1CT (Average) 13.64 7.33 615 492.0 27 10.0 28 193 1.9 12.9 0.135 0.154 4.7 5.4 1.2 2.7 3.6 2.8 0.9 4.3 3.8 1.0

Calculated Heads 186.02 100.00 952 670 158 27 58 508 17.3 22.1 0.259 1.10 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 1186 633 153 36 128 529 16.6 20.6 0.263 1.15 Accountabilities 80.3% 105.8% 103% 74.8% 45.4% 96.1% 104.7% 107.2% 98.5% 95.7%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 144.57 77.72 828 622 181 28 59 534 20.2 23.1 0.268 1.251 67.6 72.2 89.3 80.8 78.8 81.6 90.7 81.2 80.4 81.2 1CC1 - 1CC2 164.87 88.63 960 676 173 29 61 537.5 19.1 23.0 0.271 1.203 89.4 89.5 97.1 94.6 93.1 93.7 97.4 92.0 92.9 92.0 1CC1 - 1CC3 170.49 91.65 976 683 169 28 61 535 18.7 22.9 0.270 1.182 94.0 93.5 98.4 97.0 95.7 96.5 98.8 94.9 95.5 98.5 1CC1 - 1CC4 172.38 92.67 979 684 168 28 61 533 18.6 22.8 0.269 1.174 95.3 94.6 98.8 97.3 96.4 97.2 99.1 95.7 96.2 99.0

113

Table D-21: Flotation Data Sheet Z10-10-2

Composite: Plant Rougher Concentrate Date: 20-Dec-12 Z10 10-2 Plant Sample from Feb 10, 2009 Purpose: Repeat of Baseline Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added Flotation Time (min) Pt Conc Cell TX Cytec pH Lime W22C H55 Stage(s) 15155 3894 Grind Cond. Float mv wt (g) liters Regrind 0.0 7.2 206 Cond 5.0 Cond 1CC1 5.0 7.6 179 2.3 1CC2 5.0 7.4 250 2.3 1CC3 5.0 7.4 256 2.3 1CC4 5.0 7.4 271 2.3

Total 0.0 0.0 0.000 0.000 0.0 5.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 89.31 43.01 6308 705 389 38 120 389 18.9 37.0 0.329 3.25 16.4 32.7 59.3 27.5 42.1 36.1 60.1 50.3 33.5 43.7 1CC2 23.86 11.49 11663 1170 327 61 143 503 12.5 36.8 0.557 3.143 8.1 14.5 13.3 11.8 13.4 12.5 10.7 13.4 15.2 11.3 1CC3 14.71 7.08 15334 1317 281 76 164 543 11.9 35.5 0.568 3.364 6.6 10.1 7.1 9.1 9.5 8.3 6.2 8.0 9.5 7.5 1CC4 7.42 3.57 17023 1300 279 75 151 550 11.7 35.2 0.537 3.657 3.7 5.0 3.5 4.5 4.4 4.2 3.1 4.0 4.5 4.1

1CTA 30785 1007 136 77 105 511 7.8 21.8 0.452 3.022 1CTB 31053 1007 137 84 111 525 7.6 22.3 0.449 3.110

1CT (Average) 72.33 34.84 30919 1006.9 137 80.5 108 518 7.7 22.1 0.450 3.066 65.2 37.7 16.8 47.1 30.6 38.9 19.9 24.3 37.3 33.4

Calculated Heads 207.63 100.00 16519 928 282 59 123 463 13.5 31.6 0.422 3.20 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 87.7% 102.8% 93% 78.2% 78.6% 86.9% 95.5% 103.2% 93.8% 103.1%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 89.31 43.01 6308 705 389 38 120 389 18.9 37.0 0.329 3.248 16.4 32.7 59.3 27.5 42.1 36.1 60.1 50.3 33.5 50.3 1CC1 - 1CC2 113.17 54.51 7437 803 376 43 125 412.6 17.5 36.9 0.377 3.225 24.5 47.2 72.6 39.3 55.5 48.6 70.8 63.7 48.7 63.7 1CC1 - 1CC3 127.88 61.59 8345 862 365 47 129 428 16.9 36.8 0.399 3.241 31.1 57.3 79.7 48.4 65.0 56.9 77.0 71.7 58.2 62.5 1CC1 - 1CC4 135.30 65.16 8821 886 360 48 131 434 16.6 36.7 0.407 3.264 34.8 62.3 83.2 52.9 69.4 61.1 80.1 75.7 62.7 66.6

114

Table D-22: Flotation Data Sheet Z10-10-3

Composite: Plant Rougher Concentrate Date: 20-Dec-12 Z10 10-3 Plant Sample from Feb 10, 2009 Purpose: Reagent Screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added Flotation Time (min) Pt Conc Cell TX Cytec pH Lime W22C H55 Stage(s) 15155 3894 Grind Cond. Float mv wt (g) liters Regrind 0.0 7.8 206 Cond Pre-aeration 30.0 7.8 206 Cond 1CC1 5.0 7.6 250 2.3 1CC2 5.0 7.6 198 2.3 1CC3 5.0 7.6 258 2.3 1CC4 5.0 7.6 277 2.3

Total 0.0 0.0 0.000 0.000 0.0 30.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 76.34 36.97 5495 851 457 42 133 456 20.3 35.0 0.459 3.69 12.5 34.1 59.2 25.8 40.7 36.6 56.5 41.4 42.5 44.4 1CC2 28.11 13.61 10593 1075 305 60 131 491 13.5 37.9 0.460 2.430 8.9 15.9 14.6 13.6 14.8 14.5 13.9 16.5 15.7 10.8 1CC3 14.98 7.25 12766 1086 252 57 135 486 11.6 36.9 0.407 3.105 5.7 8.5 6.4 6.9 8.1 7.6 6.4 8.6 7.4 7.3 1CC4 11.90 5.76 15824 1066 225 63 127 514 10.2 34.9 0.407 3.266 5.6 6.7 4.5 6.0 6.1 6.4 4.4 6.4 5.9 6.1

1CTA 30282 901 119 77 102 446 6.8 23.2 0.374 2.637 1CTB 29849 866 120 81 99 438 6.8 23.2 0.250 2.668

1CT (Average) 75.18 36.41 30066 883.7 120 79.0 101 442 6.8 23.2 0.312 2.652 67.3 34.8 15.3 47.7 30.3 34.9 18.8 27.1 28.5 31.4

Calculated Heads 206.51 100.00 16256 923 285 60 121 461 13.2 31.2 0.399 3.07 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 86.3% 102.2% 94% 79.2% 77.4% 86.5% 93.6% 102.0% 88.7% 99.1%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 76.34 36.97 5495 851 457 42 133 456 20.3 35.0 0.459 3.689 12.5 34.1 59.2 25.8 40.7 36.6 56.5 41.4 42.5 41.4 1CC1 - 1CC2 104.45 50.58 6867 911 416 47 132 465.4 18.5 35.8 0.459 3.350 21.4 50.0 73.8 39.4 55.5 51.1 70.4 57.9 58.2 57.9 1CC1 - 1CC3 119.43 57.83 7607 933 396 48 133 468 17.6 35.9 0.453 3.319 27.1 58.5 80.2 46.3 63.6 58.7 76.8 66.5 65.6 62.5 1CC1 - 1CC4 131.33 63.59 8351 945 380 49 132 472 16.9 35.8 0.448 3.314 32.7 65.2 84.7 52.3 69.7 65.1 81.2 72.9 71.5 68.6

115

Table D-23: Flotation Data Sheet Z10-10-4

Composite: Plant Rougher Concentrate Date: 20-Dec-12 Z10 10-4 Plant Sample from Feb 10, 2009 Purpose: Reagent Screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime SO W22C H55 Stage(s) 7262 3894 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 7.0 294 Cond 2g Pre-aeration with SO2 30.0 3.3 203

Cond 2.76g 2 mls 2 µL No Air or SO2 10.0 11.9 68 1CC1 5.0 11.6 -47 2.3

1CC2 2 mls 2 µL 5.0 11.5 -37 2.3 1CC3 2 mls 5.0 11.5 -30 2.3 1CC4 2 mls 5.0 11.2 -15 2.3

Total 0.0 0.0 0.000 0.000 0.0 40.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 29.32 14.05 3173 331 173 10 85 395 28.2 27.8 0.142 1.84 2.8 4.9 10.6 2.3 11.2 12.7 30.2 13.4 4.7 8.8 1CC2 20.97 10.05 4588 528 361 24 90 446 25.2 27.2 0.191 3.467 2.9 5.6 15.9 4.0 8.5 10.2 19.3 9.4 4.5 11.8 1CC3 4.65 2.23 10677 1001 325 56 102 517 17.3 24.7 0.464 4.130 1.5 2.3 3.2 2.1 2.1 2.6 2.9 1.9 2.4 3.1 1CC4 3.08 1.48 14128 1235 255 69 106 551 12.9 23.7 0.558 3.946 1.3 1.9 1.6 1.7 1.5 1.9 1.4 1.2 1.9 2.0

1CTA 20211 1134 219 77 115 444 8.6 30.6 0.524 3.100 1CTB 20065 1124 216 73 112 436 8.2 29.3 0.503 2.962

1CT (Average) 150.65 72.20 20138 1129.0 218 75.0 114 440 8.4 29.9 0.514 3.031 91.5 85.3 68.7 89.9 76.7 72.6 46.2 74.1 86.5 74.3

Calculated Heads 208.67 100.00 15892 955 229 60 107 438 13.1 29.1 0.428 2.95 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 84.3% 105.8% 75% 79.3% 68.4% 82.1% 92.7% 95.3% 95.1% 95.0%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 29.32 14.05 3173 331 173 10 85 395 28.2 27.8 0.142 1.841 2.8 4.9 10.6 2.3 11.2 12.7 30.2 13.4 4.7 13.4 1CC1 - 1CC2 50.29 24.10 3763 413 251 16 87 416.4 26.9 27.5 0.162 2.519 5.7 10.5 26.5 6.3 19.7 22.9 49.5 22.8 9.2 22.8 1CC1 - 1CC3 54.94 26.33 4348 463 258 19 88 425 26.1 27.3 0.188 2.655 7.2 12.8 29.7 8.4 21.8 25.5 52.4 24.7 11.6 23.7 1CC1 - 1CC4 58.02 27.80 4867 504 257 22 89 432 25.4 27.1 0.207 2.724 8.5 14.7 31.3 10.1 23.3 27.4 53.8 25.9 13.5 25.7

116

Table D-24: Flotation Data Sheet Z10-10-5

Composite: Plant Rougher Concentrate Date: 20-Dec-12 Z10 10-5 Plant Sample from Feb 10, 2009 Purpose: Reagent Screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime SO W22C H55 Stage(s) 7262 3894 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 7.4 131 Cond 2g Pre-aeration with SO2 30.0 4.1 152 Cond 1CC1 5.0 4.9 123 2.3 1CC2 5.0 5.0 122 2.3 1CC3 5.0 5.0 123 2.3 1CC4 5.0 5.2 119 2.3

Total 0.0 0.0 0.000 0.000 0.0 30.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 8.01 3.93 12295 1144 235 54 97 489 11.3 21.1 0.537 3.59 3.0 4.6 4.0 3.5 3.6 4.3 3.4 2.9 4.7 4.9 1CC2 4.11 2.02 14494 1282 250 76 113 557 12.4 23.0 0.597 3.841 1.8 2.7 2.2 2.5 2.2 2.5 1.9 1.6 2.7 2.7 1CC3 3.85 1.89 14878 1295 249 74 107 555 12.3 22.9 0.597 3.796 1.7 2.5 2.0 2.3 1.9 2.3 1.8 1.5 2.5 2.5 1CC4 5.32 2.61 15517 1326 246 72 107 547 12.6 23.8 0.606 3.951 2.5 3.5 2.8 3.1 2.7 3.2 2.5 2.1 3.5 3.6

1CTA 16555 949 228 59 104 442 13.0 30.0 0.438 2.818 1CTB 16275 938 228 60 105 442 13.2 29.5 0.430 2.779

1CT (Average) 182.51 89.55 16415 943.9 228 59.5 105 442 13.1 29.7 0.434 2.799 91.0 86.7 89.0 88.6 89.6 87.7 90.4 91.9 86.6 86.3

Calculated Heads 203.80 100.00 16162 975 230 60 104 451 13.0 29.0 0.449 2.90 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 85.8% 108.0% 76% 79.2% 67.0% 84.6% 91.8% 94.7% 99.8% 93.5%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 8.01 3.93 12295 1144 235 54 97 489 11.3 21.1 0.537 3.595 3.0 4.6 4.0 3.5 3.6 4.3 3.4 2.9 4.7 2.9 1CC1 - 1CC2 12.12 5.95 13041 1191 240 61 102 512.3 11.6 21.7 0.558 3.678 4.8 7.3 6.2 6.0 5.8 6.8 5.3 4.5 7.4 4.5 1CC1 - 1CC3 15.97 7.84 13484 1216 242 64 104 523 11.8 22.0 0.567 3.707 6.5 9.8 8.2 8.3 7.7 9.1 7.1 6.0 9.9 10.1 1CC1 - 1CC4 21.29 10.45 13992 1243 243 66 104 529 12.0 22.5 0.577 3.768 9.0 13.3 11.0 11.4 10.4 12.3 9.6 8.1 13.4 13.7

117

Table D-25: Flotation Data Sheet Z10-10-6

Composite: Plant Rougher Concentrate Date: 20-Dec-12 Z10 10-6 Plant Sample from Feb 10, 2009 Purpose: Reagent Screening Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime SO W22C H55 Stage(s) 7261A 3894 2 Grind Cond. Float mv wt (g) liters Regrind 0.0 7.4 129 Cond 2g Pre-aeration with SO2 30.0 4.2 152

Cond 1.20g 3 mls 2 µL 10.0 11.3 -124 1CC1 5.0 10.6 4 2.3

1CC2 2 mls 2 µL 5.0 10.1 28 2.3 1CC3 2 mls 5.0 9.7 37 2.3 1CC4 2 mls 5.0 9.4 62 2.3

Total 0.0 0.0 0.000 0.000 0.0 40.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 8.97 4.33 12929 1161 242 64 100 502 12.5 22.2 0.554 3.59 3.5 5.4 4.7 4.3 4.1 5.1 4.0 3.2 5.3 5.2 1CC2 5.17 2.50 11810 1076 253 53 99 516 16.1 24.1 0.509 3.583 1.8 2.9 2.8 2.1 2.3 3.0 3.0 2.0 2.8 3.0 1CC3 4.63 2.24 13662 1213 262 67 106 538 14.4 24.2 0.565 3.882 1.9 2.9 2.6 2.3 2.2 2.8 2.4 1.8 2.8 2.9 1CC4 3.74 1.81 14951 1253 254 65 105 538 13.8 25.3 0.591 4.082 1.7 2.4 2.0 1.8 1.8 2.3 1.9 1.5 2.4 2.5

1CTA 16621 895 222 63 103 428 12.8 29.4 0.432 2.816 1CTB 16574 221 65 110 412 13.9 32.5 0.452 2.947

1CT (Average) 184.43 89.12 16598 895.3 222 64.0 107 420 13.4 31.0 0.442 2.882 91.1 86.4 87.9 89.5 89.6 86.8 88.7 91.5 86.7 86.4

Calculated Heads 206.94 100.00 16223 925 225 64 106 431 13.4 30.2 0.454 2.97 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 86.1% 102.4% 74% 84.0% 67.9% 80.8% 95.0% 98.6% 100.9% 95.9%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 8.97 4.33 12929 1161 242 64 100 502 12.5 22.2 0.554 3.589 3.5 5.4 4.7 4.3 4.1 5.1 4.0 3.2 5.3 3.2 1CC1 - 1CC2 14.14 6.83 12520 1130 246 60 100 507.3 13.8 22.9 0.537 3.587 5.3 8.3 7.5 6.4 6.4 8.1 7.0 5.2 8.1 5.2 1CC1 - 1CC3 18.77 9.07 12802 1150 250 62 101 515 13.9 23.2 0.544 3.660 7.2 11.2 10.1 8.7 8.6 10.9 9.4 7.0 10.9 11.1 1CC1 - 1CC4 22.51 10.88 13159 1167 251 62 102 519 13.9 23.6 0.552 3.730 8.9 13.6 12.1 10.5 10.4 13.2 11.3 8.5 13.3 13.6

118

Table D-26: Flotation Data Sheet Z10-10-7

Composite: Plant Rougher Concentrate Date: 30-Jan-13 Z10 10-7 Plant Sample from Feb 10, 2009 Purpose: Confirm Conditions Grind:~80% passing µm Sands ~80% passing µm Feed: 230 grams of as-received plant final concentrate Operator: ZZ Grind: Rod Mill C % minutes Conditions: Regrind: RMA % 0.0 minutes

Conditions Reagents Added Flotation Time (min) Pt Conc Cell Cytec Cytec pH Lime SO W22C H55 Stage(s) 7261A 3894 2 Grind Cond. Float mv wt (g) liters Conditions 0.0 7.7 101 Conditions Condition with SO2/air 23 30.0 3.1 193 Conditions 28 4 µL 4 mls 10.0 12.0 -342 Cleaner 1 1.0 2.3 Cleaner 2 2.0 12.1 -54 2.3 Cleaner 3 4 µL 4 mls 3.0 5.0 11.9 -59 2.3 Cleaner 4 4mls 5.0 11.7 -31 2.3 Cleaner 5 4mls 7.0 11.57 -4

Total 28.4 0.0 0.000 0.000 0.0 43.0 20.0

Metallurgical Results Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 6.59 1.57 12054 1053 223 76 71 15 14.2 24.5 0.466 2.71 1.1 1.7 1.5 1.5 1.6 1.7 1.5 1.2 1.7 1.7 1CC2 18.18 4.34 10801 917 196 71 59 15 19.6 27.8 0.402 2.47 2.8 4.0 3.6 3.9 3.7 4.8 5.9 3.7 4.0 4.4 1CC3 66.01 15.74 4631 464 328 36 46 14 28.3 28.1 0.159 2.62 4.4 7.4 22.1 7.2 10.4 16.6 30.9 13.5 5.8 16.9 1CC4 13.76 3.28 11644 1069 277 72 61 16 18.4 27.8 0.478 3.17 2.3 3.6 3.9 3.0 2.9 3.9 4.2 2.8 3.6 4.3 1CC5 11.18 2.67 14393 1285 249 89 72 18 14.4 26.1 0.535 3.24 2.3 3.5 2.8 3.0 2.7 3.4 2.7 2.1 3.3 3.5 0.0 0.0 1CTA 20166 1082 213 89 76 13 10.9 34.9 0.488 2.33

1CT (Average) 303.60 72.40 20166 1082.4 213 89.0 76 13 10.9 34.9 0.488 2.33 87.1 79.8 66.1 81.4 78.7 69.6 54.8 76.7 81.6 69.2

Calculated Heads 419.32 100.00 16753 982 234 79 70 14 14.4 32.9 0.433 2.44 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 Assayed Heads 18842 903 304 76 156 533 14.2 30.6 0.450 3.10 Accountabilities 88.9% 108.8% 76.8% 104.1% 44.8% 2.6% 101.8% 107.4% 96.2% 78.7%

Cumulative Assays & Distribution Assays Distribution, % Products Weight % As Bi Cd Sb Se Ag Cu Fe Pb Zn As Bi Cd Sb Se Ag Cu Fe Pb Zn grams % Solids ppm ppm ppm ppm ppm gpt % % % % 1CC1 6.59 1.57 12054 1053 223 76 71 15 14.2 24.5 0.466 2.71 1.1 1.7 1.5 1.5 1.6 1.7 1.5 1.2 1.7 1.2 1CC1 - 1CC2 24.77 5.91 11134 954 203 72 62 14.9 18.1 26.9 0.419 2.53 3.9 5.7 5.1 5.4 5.3 6.5 7.4 4.9 5.7 4.9 1CC1 - 1CC3 90.78 21.65 6405 598 294 46 50 15 25.5 27.8 0.230 2.60 8.3 13.1 27.2 12.6 15.7 23.1 38.3 18.4 11.5 23.0 1CC1 - 1CC5 115.72 27.60 7800 720 288 53 54 15 23.6 27.6 0.289 2.73 12.9 20.2 33.9 18.6 21.3 30.4 45.2 23.3 18.4 30.8

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