University of Nevada, Reno Optimization of a Copper

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University of Nevada, Reno

Optimization of a Copper/Molybdenum Flotation Circuit by Depressing Pyrite from

the Rougher Flotation Stage Using Different Scheme of Reagents

A thesis submitted in partial fulfillment of the

requirements for the degree of Master Of Science in

Metallurgical Engineering

Wander A. Valderrama

Dr. Carl C. Nesbitt/Thesis Advisor

May 2020

We recommend that the thesis prepared under our supervision by

entitled

be accepted in partial fulfillment of the requirements for the degree of

Advisor

Committee Member

Graduate School Representative

David W. Zeh, Ph.D., Dean Graduate School

Abstract

A study to determine methods to hinder the flotation of pyrite, other than the specific use of lime, in rougher flotation was conducted. Specifically, the goal was to determine an effective alternative to pH adjustment using lime. Experiments were conducted using the most common selection of reagents: xanthate and lime to recover copper and reject iron to establish a base test to compare other reagent schemes with.

This work included flotation reagents such as dithiophosphate, thiocarbamate and AERO

3302 promoter as chalcopyrite collectors, fuel oil as molybdenite collector, and sodium sulfite, inc sulfate, calcium hydroxide and potassium dichromate as pyrite depressants from literature review and expert suggestions. Of the reagents tested, potassium dichromate was included as a modification to simulate chrome grinding media effect.

Some studies have documented that there is a galvanic interaction between the grinding media – sulfide mineral system that affects the recovery of copper. Thus, researchers have used chrome grinding media resulting in better recoveries. Due to manufacturing issues, this study could not perform tests using chrome grinding media. Thus, it implemented the use of potassium dichromate as a depressant instead of replacing the forged steel grinding media.

Acknowledgements

The author wishes to thank Dr. Carl C. Nesbitt, Associate Professor of Metallurgical

Engineering, for his support and advice during this research work. His financial and technical support are highly appreciated.

I would also like to thank Dr. Nesbitt for convincing me to stay for two more years in graduate school and add a milestone to my professional development.

Lastly, the author wishes to thank his family and best friend that even though they are miles apart, they showed their emotional support and cheerleading skills through videocalls to get through the tough times. My school friends and my students who kept schooling interesting during my stay at School of Earth Science and Engineering:

Department of Mining and Metallurgical Engineering.

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Table of Contents

Abstract ------i

Acknowledgements ------ii

List of Tables ------iv

List of Figures ------v

Introduction ------1

Background ------1

Flotation------4

Depression of Pyrite ------6

Analytical Method ------10

Experimental Procedure ------13

Reagents Selection ------13

Specifications of Experiment ------19

Results and Discussion ------22

Further Studies ------34

Conclusions ------36

Reference ------38

Appendix ------40

List of Tables

Table 1: Schemes of reagents used to depress pyrite in a Copper-molybdenum rougher flotation circuit (*)...... 14

Table 2: AERO 3302 chemical and physical characteristics (7)...... 16

Table 3: Reagent schemes used to improve copper recovery and depress pyrite (*)...... 23

Table 4: Assay analysis of a 500-g after using ICP analytical method...... 24

Table 5: Percentage for each mineral found in the rougher concentrate corresponding to each scheme...... 25

Table 6: Recovery for copper and iron in the rougher concentrate using auto and semi qualitative settings from XRD analysis...... 26

Table 7: Selectivity ratio for metal recoveries...... 33

List of Figures

Figure 1: Global Copper production since the 1900s. ------2

Figure 2: Global Molybdenum production since the 1900s. ------3

Figure 3: Chemical structure of xanthate. ------15

Figure 4: Chemical structure of xanthate allyl ester. ------16

Figure 5: Standard scheme modifying pH (Auto-qualitative). ------28

Figure 6: Standard scheme modifying pH (Semi-qualitative). ------28

Figure 7: Comparison of Cu and Fe recovery (Auto-qualitative). ------31

Figure 8: Comparison of Cu and Fe recovery (Semi-qualitative). ------32

Figure 9: Selectivity ratio summariing metal recovery for each scheme. ------34

Figure 10:Standard scheme using xanthate, fuel oil and limestone as flotation reagents

(AQ). ------40

Figure 11: Standard scheme using xanthate, fuel oil and limestone as flotation reagents

(SQ). ------40

Figure 12: Scheme 1 using thiocarbamate, AERO 3302, fuel oil and sodium sulfite as flotation reagents (AQ). ------41

Figure 13: Scheme 1 using thiocarbamate, AERO 3302, fuel oil and sodium sulfite as flotation reagents (SQ). ------41

Figure 14: Scheme 2 using AERO 3302, fuel oil and potassium dicromate as flotation reagents (AQ). ------42

Figure 15: Scheme 2 using AERO 3302, fuel oil and potassium dichromate as flotation reagents (SQ). ------42

Figure 16: Scheme 3 using AERO 3302, fuel oil and inc sulfate as flotation reagents

(AQ). ------43

Figure 17: Scheme 3 using AERO 3302, fuel oil and inc sulfate as flotation reagents

(SQ). ------43

Figure 18: Scheme 4 using thiocarbamate, AERO 3302, fuel oil and sodium sulfite (AQ).

------44

Figure 19: Scheme 4 using thiocarbamate, AERO 3302, fuel oil and sodium sulfite as flotation reagents (SQ). ------44

Figure 20: Scheme 4.1 using thiocarbamate, AERO 3302, fuel oild and potassium dichromate as flotation reagents (AQ). ------45

Figure 21: Scheme 4.1 using thiocarbamate, AERO 3302, fuel oil and potassium dichromate as flotation reagents (SQ). ------45

Figure 22: Scheme 4.2 using thiocarbamate, AERO 3302, fuel oil and potassium dichromate as flotation reagents (AQ). ------46

Figure 23: Scheme 4.2 using thiocarbamate, AERO 302, fuel oil and potassium dichromate as flotation reagents (SQ). ------46

Figure 24: Scheme 5 using AERO 3302, fuel oil, potassium dichromate and lime as flotation reagents (AQ). ------47

Figure 25: Scheme 5 using AERO 3302, fuel oil, potassium dichromate and lime as flotation reagents (SQ). ------47

Figure 26: Scheme 6 using AERO 3302 and lime as flotation reagents (AQ). ------48

Figure 27: Scheme 6 using AERO 3302 and lime as flotation reagents (SQ). ------48

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Figure 28: Scheme 7 using AERO 3302, fuel oil, potassium dichromate, lime and sodium sulfite as flotation reagents (AQ). ------49

Figure 29: Scheme 7 using AERO 3302, fuel oil, potassium dichromate, lime, and sodium sulfite as flotation reagents (SQ). ------49

Introduction Background

The development of human civiliation and the discovery of mineral resources have been associated since the beginning of time. Obsession with metals started when early men made the switch from the stone tools used to perform farming tasks during a period known as the Neolithic Age. This era of metal-based tools was named the Brone

Age and dated from 3000 – 1200 BCE. Between the transition from stone tools to the obsession with metals, the Copper Age or Chalcolithic period was roughly developed from 4500 – 3500 BCE overlapping the Brone Age. Eastern Mediterranean regions such as the island of Cyprus mined copper from naturally occurring metallic copper. Its early uses ranged from jewelry, utensils, tools, weapons, and then gradually increasing over the years during the 20th century with mass adoption of electricity (5). Properties such as electrical and thermal conductivity, corrosive resistance, and malleability to fabricate wires, pipes and sheets have contributed to make copper a valuable base metal on the commodity market (Fig. 1).

2,500

2,000

1,500

1,000 x 10000x

millions of tonnes/year of millions 500

World mine production of copper, copper,of mine productionWorld 0 1900 1920 1940 1960 1980 2000 2020 Year

Figure 1: Global Copper production since the 1900s showing an exponential increase through the years due to its various use in different industries. The graph corresponds to data compiled in Copper Statistics by USGS (K.E. Porter, D.L. Edelstein, and M. Brininstool, Jan 2017, USGS, 2017a).

Copper ore is found in the earth’s crust as copper-iron-sulfide and copper sulfide minerals such as chalcopyrite (CuFeS2), bornite (Cu5FeS4) and chalcocite (Cu2S) (8).

Also, minerals in the oxidied upper region of carbonates, oxides, hydroxy-silicates, and sulfates are also considered copper ores. All these ore containing minerals can have low concentrations of copper ranging from 0.5% (open pit mines) to 1 or 2% (underground mines). Sulfides are recovered by concentration, smelting, and refining while oxides are usually extracted through hydrometallurgical methods (leaching solution, solvent extraction, and electrowinning circuit).

The minerology of the ore determines the extraction process implemented to recover a specific metal. In the case of copper, approximately 80% of copper-from-ore is originated in Cu-Fe-S ores (5). At low pH and/or in the presence of a bacteria, Cu-Fe-S ores can be treated through aqueous solutions yet, most of copper extraction worldwide is conducted through pyrometallurgical methods. Additionally, many copper sulfide ore

3 deposits also contain other minerals such as molybdenite as a secondary mineral, and

(among other minerals) pyrite and silicate as gangue minerals (9). Sulfide copper ores contain by-product molybdenite, a major by-product mineral from copper concentrators.

The global molybdenum production has a use of 50% as an alloying agent in the steel industry for high-speed steel to produce cutting material tools. Other uses of molybdenum include aircraft parts, some missiles, and the nuclear power industry.

Compared to base and precious metals, molybdenum usage is minor with maximum production of 268 million tonnes in 2014 (Fig.2). Most of this by-product molybdenum comes from copper sulfide ores, and, at lower recoveries from porphyry copper ores (4).

However, since 2004 molybdenum has increased its market price nearly “10-fold resulting in a renewed emphasis on production of this metal, including efforts to optimie its flotation recovery” (4). 30.0

25.0

20.0

15.0

10.0 x 10000x

5.0

millions of tonnes/year of millions 0.0 1900 1920 1940 1960 1980 2000 2020

molybdenum, of mine productionWorld Year

Figure 2: Global Molybdenum production since the 1900s showing the exponential increase starting in 2004. The graph corresponds to data compiled in Molybdenum Statistics by USGS (T.D. Kelly, M.J. Magyar, and D.E. Polyak, Jan 2017, USGS, 2017b).

Copper and molybdenum recoveries when occurring as sulfide minerals primarily rely on a process called flotation. This process was first patented by William Haynes in

1860 “who claimed that sulfides could be floated by oil and non-sulfide minerals could be removed by washing in a powdered ore” (6). The first commercial cell for a flotation plant was built by the Bessel brothers in Dresden, Germany. Along with that, Carrie

Everson patented the first flotation process of sulfide ores in 1885 (6). Bessel bothers’ commercial cell and Everson’s patent represented an important milestone in the history of flotation as a process to separate valuable minerals from gangue material. Developments of the process continued from “being a research topic in the lab to a more commercially valuable tool, in the early [20th] century” (6). Thus, froth flotation was implemented in the United States for the first time in 1911 by James M. Hyde in Basin, Montana. The flotation circuit was able to remove entrained gangue particles from concentrates, and it positioned flotation as the process to recover valuable minerals from sulfide ores. From there, flotation has been “the most important and versatile industrial process for the separation and concentration of minerals” (6).

Flotation

Flotation, as a process is defined as the separation of “finely divided particles from each other by means of using air bubbles under hydrodynamic environment” (6).

Froth flotation entails crushing and grinding of the ore to obtain a proper liberation of particles to apply flotation principles. The process is conducted in water from the grinding circuit because it relies on the physical properties of the particles and water.

Hydrophilic particles (hydrophile – water-wettable) sink in water without the necessity of changing their surface chemistry while hydrophobic particles (hydrophobe – water-

5 repellant) repel water and easily float on top of a solid-liquid mixture. The challenge arises when the valuable mineral is hydrophilic along with the gangue material or the gangue material attaches to the valuable mineral which does not allow the recovery of valuable particles solely. To change that, chemical reagents are required for this process.

Collectors, activators, depressants, modifiers and frothers are specific chemicals with the purpose of affecting particles and the water properties so that flotation can happen and obtain an enriched concentrate with valuable metals for further refining processes.

Most copper sulfide ores depend on froth flotation for the recovery of copper along with any other profitable by-product. According to the principles of froth flotation,

• sulfide minerals are hydrophilic particles but by conditioning their surface

chemistry with chemical reagents (collectors), they become water repellent.

• this repellant nature is selectively applied to Cu minerals leaving the gangue

material wetted

• rising air bubbles collide with the ‘now-water repellent Cu minerals’ causing

Cu mineral particles to float

• wetted mineral particles, as they are not bound to air, stay sunken in water (5)

Similarly, molybdenite minerals can be floated by applying flotation principles.

However, these minerals are naturally hydrophobic or water repellent which facilitates the separation process along with other valuable minerals from gangue material (9). Its hydrophobicity allows molybdenite to easily float “among other sulfide minerals, but its enrichment or concentration by flotation is not always simple” (9). Factors such as mineralogy, slime coatings, optimiation of copper metallurgy at the expense of

6 molybdenum, grinding and liberation, and flotation reagents might affect molybdenum recovery from porphyry copper ores (4). Although molybdenite floats without major reagent conditioning, the mentioned factors tend to impact the enrichment of the mineral by flotation. Thus, ores containing molybdenite as recoverable material and clay as gangue material result “in the decrease of the recovery of molybdenum to just below 40% recovery” (9). Clay has a significant impact on the flotation recovery of copper and molybdenum because the molybdenite flotation is negatively affected by surface oxidation, and calcium ions revealed an affinity to absorb onto molybdenite surface causing the reverse of its eta potential. Due to this interaction, molybdenite floatability decreases resulting in low recoveries from porphyry ores (4). To mitigate this effect caused by clay, a study from 2008 found that by decreasing the feed density going to flotation, molybdenite improved its recovery while copper remained unchanged or improved marginally (4). Nonetheless, the objective of this thesis work is mainly focused on chalcopyrite and pyrite but understanding that molybdenite occurs as a byproduct of the process.

Depression of Pyrite

As previously mentioned, the main impurity found in the recovery process of copper as chalcopyrite is pyrite. From a mineralogical perspective, pyrite (FeS2) or iron disulfide is the most abundant sulfide mineral on earth, but also the most common gangue material when recovering other base metal minerals (3, 10). Elevated amounts of pyrite degrade mineral value of concentrates and increases iron and sulfur content which escalates the smelting costs (3, 11). Valuable minerals such as sphalerite, chalcopyrite,

7 galena, gold as well as coal are commonly found with pyrite (11). Yet, pyrite has “a low economic significance” (11) when compared to other sulfide minerals, and it becomes a waste or gangue material in many ore deposits. Additionally, pyrite presence affects further pyrometallurgical processes. For instance, it is the major sulfur constituent contributing to high-sulfur coals. Its combustion from pyrite-bearing coal will release sulfur components – SOx (SO, SO2 or SO3), which has been attributed to causing acid rain (11). Thus, economic, and environmental aspects urge the separation of pyrite from valuable minerals at early stages of processing. Moreover, froth flotation achieves this goal through the addition of reagents to change the surface chemistry of the slurry

- containing valuable and gangue minerals. Xanthate (ROCSS 2M) is commonly used as the collector reagent in the flotation of base metal sulfide minerals. The absorption of this collector on sulfide mineral surfaces involves electrochemical reactions including charge transfer in the chemisorption of xanthate (Reaction (1)), oxidation of xanthate to dixanthogen (Reaction (2)) and/ or the formation of metal xanthate (Reaction (3)) (11).

− − 푋 → 푋푎푑푠 + 푒 (1)

− − 2푋 ↔ 푋2 + 2푒 (2)

− 푀푆 + 2푋 → 푀푋2 + 푆 + 2푒 (3)

- X , Xads, X2, MS, MX2 and S represent xanthate ions, the adsorbed xanthate, dixanthogen, sulfide mineral, metal xanthate, and elemental sulfur or polysulfide, respectively. From the agents involved in the flotation process, dixanthogen is considered “the predominant species responsible for the hydrophobicity of pyrite” (11). Hence, these electrochemical

8 reactions and pyrite flotation have shown a strong correlation causing various phenomena such as “changes in pyrite surface chemistry due to oxidation, the interaction of pyrite with other components, adsorption of collector and precipitation of metals on surface”

(3).

Interactions at an electrochemical level between pyrite, reagents and valuable minerals, such as chalcopyrite, start in the grinding circuit. The beneficiation of copper sulfide ores involves wet milling circuits using ball mills and, in some cases, reagents are added at this early stage of the recovery process. Conditions such as pH, pulp oxidation potential, grinding media, galvanic interactions or electrochemical reactions among others can be controlled at any stage of the processing circuit. Nonetheless, all the above conditions have shown major potential affecting the floatability of copper minerals. As a result, the process can either float more pyrite as impurity in the concentrate or depress valuable mineral resulting in low recoveries.

In copper flotation circuits, chalcopyrite is known to have better recoveries when xanthate is added to the system; however, it increases the floatability of pyrite. Previously stated, the addition of xanthate causes a chain of reactions (see Eq 1-3) which can produce dixanthogen as a product. Dixanthogen has “been known as the main hydrophobic produc[t] formed on the pyrite surface in the presence of xanthates” (3).

Yet, studies conducted by Trahar shows that xanthate as collector indeed floats pyrite at higher cathodic potentials than the equilibrium potentials of xanthate/dixanthogen couples (3). Moreover, dixanthogens are not solely responsible for floating pyrite in a copper concentrate suggesting other mechanisms may be present. Another agent causing this phenomenon is “the formation of the ferric xanthate compounds and absorption of

9 ethyl xanthate ions may be the reason why the pyrite flotation occurs at more cathodic potentials than the potential required for the dixanthogens formation from xanthates” (3).

Reducing pyrite floatability is achievable by modifying the pH to alkaline conditions causing its depression. Pyrite tends to depress at high pH ranging from 9 to 11 for rougher stages and pH>11 for cleaning stages. Such conditions are obtained by adding lime (CaO) in its slurry form (Ca(OH)2) to the flotation circuit. In alkaline conditions,

“the interaction between hydroxyl ions and surface species results in hydrophilic iron hydroxyl on pyrite surface preventing the absorption of [collectors]” (11). Increasing the pH conditions oxidizes the pyrite surface suggesting that “the oxidation of pyrite replace[s] the oxidation of xanthate” (11). This electrochemical interaction between pyrite, xanthate and pH modifiers inhibits the formation of dixanthogen at high pH. Thus, under alkaline conditions, pyrite is more cathodic than chalcopyrite causing the surface oxidation to form on pyrite which leads to its depression (11). However, while pH adjustment is a known method for the depression of pyrite in copper and molybdenite flotation circuits, it is not cheap. Some large mines (such as the new copper project in

Panama: Cobre Panama) can process thousands of tons of ore per hour. The water volume required to slurry this much ore can be a tremendous volume. Trying to raise the pH of millions of gallons of fresh water would be a huge cost on the process. So alternative processes are needed to mitigate the consumables cost.

Evaluation and selection for reagent schemes were based on literature reviews and suggestions by experts on the subject (refer Sec. Reagents Selection). The primary objective of this reagent’s selection is to observe what combination of chemicals or single chemical would depress pyrite from the rougher circuit. The thought is, if pyrite can be

10 depressed within the first separation process, it will be less troublesome for the cleaning circuit and molybdenite circuits, as well. To compare the results of the proposed reagents, the chosen standard metric was based on pH modifications since it is the most common method to depress pyrite from a chalcopyrite-pyrite pulp. This phenomenon is caused by the addition of lime or calcium hydroxide – CaO, Ca(OH)2. Ca(OH)2 or limestone –

CaCO3. These calcium sources has been the principal reagent to recover a concentrate of copper containing chalcopyrite and pyrite. By modifying the pulp pH to alkaline conditions, pyrite particles tend to depress by becoming more hydrophilic due to “the relative abundance of the hydrophilic hydroxide to the hydrophobic sulfur species” (3).

This ratio between hydrophilic hydroxide and hydrophobic sulfur species largely depends on the solubility of the oxidation products. Experiments have demonstrated that “the pyrite surface is oxidied from the mineral lattice to the surface [because of] their high mobility” (3).

Analytical Method

Full chemical analysis (by ICP or acid digestion) would not be useful for this project. As the project’s objective was to reject pyrite, an analytical process that fuses all the iron into solution would not be able to delineate between the iron from the pyrite, and other minerals (especially chalcopyrite). Thus, a technique was needed that could determine the relative compositions by mineral species, not chemical analysis. For this project, X-ray diffraction (XRD) presented the best analysis results possible. Rougher concentrate from each scheme of reagents was subject to an analytical study to determine what minerals are present within the collected sample. For minerals with variable

11 formulas and structures, XRD analysis was the best method for identifying and determining the proportion within the sample.

The experimental science behind this technique is known as x-ray crystallography in which the crystalline structure causes a beam of incident X-rays to diffract into many directions. The specific angles at which the crystal diffract the beam into the detector correspond to planes of the crystal. Each crystal has a characteristic pattern of diffraction angles and corresponding intensity of the diffracted beam (13). When obtaining the readings from an XRD analysis, it shows a graph with different peaks at various intensities. Those peaks in the data were compared to peaks in mineral scans from a digital library to identify the minerals present. The process is typically only qualitative— it indicates if the minerals are in the sample, or not. However, the instrument used in this study was a Bruker Phaser 2. This instrument can compare relative counts of the different minerals identified and provide a “semi-quantitative” analysis of the different minerals present. While this was not an accurate or precise measure, the relative percentages were an indicative comparison of the success of the reagent schemes. This quantitative data allowed for further numerical analysis (such as metal recovery) to observe the effect of a reagent to float or depress material.

Although rougher concentrates were not analyed using ICP or acid digestion, a

500-g sample of ore was sent off for a full chemical analysis to a private laboratory. The objective for this sample was to detect and measure the elements present in the ore sample. Thus, ICP (Inductively Couple Plasma) Spectroscopy was used to analye the ore

12 in study. This analytical method was based on the ioniation of a sample by an extremely hot plasma, usually made from argon gas.

Experimental Procedure

The general purpose of this research was to determine methods to improve the rejection of pyrite from a flotation circuit composed of a rougher and cleaner stage. The ore coming from as crushing-grinding circuit predominantly contains copper and molybdenum with gold and silver as byproducts. Due to distance issues and the fact that the project in Panama was not in operation yet an ore that had similar characteristics was used. Samples were collected from an abandoned project located in Tonopah, Nevada.

Since the main goal was to reject pyrite, the ore from Tonopah met the purpose of this research project. By implementing numerous reagent schemes, it was intended to improve molybdenite (MoS2) and chalcopyrite (CuFe2S) recovery, while rejecting most of the pyrite (FeS2) impurities possibly found in the concentrates.

Reagents Selection

These reagent schemes have been used by the industry to accomplish the goal of improving copper and molybdenum recovery while minimiing iron recovery. To compare the results from these schemes, a standard reagent, lime (CaO), was used to depress pyrite in base line testing. Considering this reagent among the schemes presented above helps to have a standard result for recovery. This comparison allowed for comparisons of which schemes gave the most cost-effective result by giving clean rougher concentrates with small amounts of pyrite impurities in it.

Table 1: Schemes of reagents used to depress pyrite in a Copper-molybdenum rougher flotation circuit (*).

Reagent Standard Dosage Scheme #1 Dosage Scheme #2 Dosage Scheme #3 Dosage Scheme #4 Dosage

CuFeS2 collector #1 Xanthate 0.25 g AERO 3302 9.00 g Dithiophosphate 0.50 g AERO 3302 9.00 g Thiocarbamate 0.50 g

CuFeS2 collector #2 ~ ~ ~ AERO 3302 4.50 g AERO 3302 4.50 g

MoS2 collector Fuel Oil 0.15 g Fuel Oil 0.15 g Fuel Oil 0.15 g Fuel Oil 0.15 g Fuel Oil 0.15 g

FeS2 depressant #1 Ca(OH)2 0.3 g Na2SO3 0.16 g K2Cr2O7 0.015 g ZnSO4 0.25 g Na2SO3 0.16 g

FeS2 depressant #2

SiO2 depressant Frother MIBC 0.20 g MIBC 0.15 g MIBC 0.10 g MIBC 0.15 g MIBC 0.15 g

Reagent Scheme #4.1 Dosage Scheme #4.2 Dosage Scheme #5 Dosage Scheme #6 Dosage Scheme #7 Dosage

CuFeS2 collector #1 Thiocarbamate 0.50 g Thiocarbamate 0.50 g AERO 3302 9.00 g AERO 3302 9.00 g AERO 3302 9.00 g

CuFeS2 collector #2 AERO 3302 2.25 g AERO 3302 0.75 g

MoS2 collector Fuel Oil 0.15 g Fuel Oil 0.15 g Fuel Oil 0.30 g Fuel Oil 0.15 g

FeS2 depressant #1 K2Cr2O7 0.015 g K2Cr2O7 0.015 g K2Cr2O7 0.015 g Ca(OH)2 0.3 g K2Cr2O7 0.015 g

FeS2 depressant #2 Ca(OH)2 0.30 g Ca(OH)2 0.30 g

SiO2 depressant Na2SO3 0.16 g Frother MIBC 0.10 g MIBC 0.20 g MIBC 0.15 g MIBC 0.20 g MIBC 0.15 g *Dosages are based on grams of reagents to grams of ore. For this work, a sample of 500 grams was used for each scheme.

Table 1 introduces the schemes of reagents that were tested to observe the depression of pyrite within each rougher concentrate. Each reagent has been commonly used to improve the recovery of copper in a rougher flotation circuit by means of pyrite depression. Lime or CaO is the preferred reagent to control the pH and therefore,

- preventing pyrite from floating along with copper particles. Xanthate (ROCSS 2M) is well-known as a copper collector used in the mining industry made from alcohols containing from 2 to 6 carbon atoms (Fig. 3). Although the collector selection is typically based on the mineral form (sulfide, oxidied and/or metallic species), their association with each other and the gangue minerals must be also considered. For this research work, xanthate was implemented in the early stages, but it was not yielding the expected results.

The mineral ore being treated was composed of with molybdenite and chalcopyrite representing valuable material with gangue minerals including pyrite and clays. The interaction between copper and gangue particles leads to gangue particles coating valuable particles. This phenomenon does not allow the collector to react with the copper particles, hence, a low recovery was obtained. To mitigate the issue, other xanthate derivatives were considered to attain a higher recovery of copper.

Figure 3: Chemical structure of xanthate. This type of collector was first introduced in 1925 and is widely used on easy-to-treat ores where selectivity (especially against iron sulfides and penalty elements) is not an issue.

AERO 3302 promoter is a copper collector deriving from and having similar properties as xanthate (Fig. 4). Yet, it is a water-insoluble and oily chemical belonging to the chemical species xanthate allyl esters. This collector is more selective than xanthate and can be used in wider pH ranges than common xanthate. Its insolubility in water requires more attention to factors such as point of addition and conditioning time, however. Xanthate allyl esters are among the most selective of all available sulfide collectors (7).

Figure 4: Chemical structure of xanthate allyl ester. Due to its long carbon chains, AERO 3302 is more selective when used to float a concentrate with less impurities.

Table 2: AERO 3302 chemical and physical properties provides detailed report of the collector’s effect on the flotation of copper for the given mineral in study (7).

Table 2: AERO 3302 chemical and physical characteristics (7).

AERO 3302 Chemical and Physical Characteristics 1. Oily collector, not soluble in water, therefore, 4. Often increase the recovery of gold and usually fed to ball mills silver 2. Effective copper collector in both alkaline and acid circuit. Also good for inc flotation in lime 5. Used for flotation of sulfidied copper- circuit. Usually used in conjunction with xanthate. oxides minerals Very selective against pyrite 3. Excellent collector for molybdenite and is, 6. Improves selective recovery of therefore, often used on copper/molybdenite ores platinum group metals

Normally, when working with an ore containing chalcopyrite as the valuable material and pyrite as the gangue material, xanthate and lime are used to recover chalcopyrite and reject pyrite. In an ideal scenario, the addition of lime to the flotation helps to control pH and, therefore, pyrite is rejected due to the interaction between hydroxyl ions and surface species causing the attachment of hydroxyl ions on pyrite surface leading to the depression effect (11). However, xanthate addition enhances the floatability of pyrite because of electrochemical reactions such as charge transfer in the chemisorption of xanthate, oxidation of xanthate to dixanthogen and/or the formation of metal xanthate (11). Particularly, dixanthogen is considered the hydrophobic specie attaching to pyrite surfaces and leading to its floatation at high cathodic potentials. These potentials overcome the equilibrium potentials of xanthate/dixanthogen couples indicating that there are other factors involved in the activation of pyrite when xanthate is added.

The mechanism of copper activation of pyrite has been documented to be responsible for pyrite flotation in alkaline conditions. Cu (OH)2 is absorbed or precipitated on the pyrite surface; then Cu (II) is reduced to Cu(I) forming Cu(I)-S together with the oxidation of sulfide (12). As a result, any excess of copper concentration remains as copper hydroxides on the pyrite surface which strongly reacts with xanthate (X) forming Cu(I)X species (See Reactions 4 and 5) (12). This reaction of copper and xanthate finally leads to the formation of dixanthogen (X2).

1 푂 + 퐻 푂 + 2푒− <=> 2푂퐻−, 2 2 2

− − 푆푠푢푟푓 − 퐶푢(퐼) + 푋 <=> 푆푠푢푟푓 − 퐶푢(퐼)푋 + 푒 (4)

1 퐶푢(푂퐻) + 2푋− <=> 퐶푢(퐼)푋 + 푋 + 2푂퐻− (5) 2,푠푢푟푓 푠푢푟푓 2 2

Many of these reactions tend to happen because of the grinding environment due to “the electrochemical potential at the mineral/solution interface and the nature of reactions occurring in a mill is considerably influenced by the type of grinding media”

(3). From an electrochemical approach, grinding creates a galvanic interaction between minerals and active grinding media (balls, rods, and liners) resulting “not only an increase in the corrosive wear of the grinding media but also a change of the surface properties and pulp chemistry of the ground minerals, which in turn affect subsequent flotation” (2). When the slurry is transferred to the flotation circuit, those interactions may have modified the chemistry causing the activation of pyrite due to the dissolution of copper species from copper sulfide minerals during grinding. Simultaneously, the iron sulfide mineral (pyrite) has modified its surface by absorbing Cu+ due to its highest rest potential compared to chalcopyrite and galena, thus leading to the absorption of xanthate into the pyrite surface. Additionally, dissolved iron ions from the grinding media react with hydroxyl species forming “hydroxyl complexes of iron which are hydrophilic and have seen as coatings on the sulfide mineral surfaces” (1). This coating on the sulfide mineral surfaces affects the floatability of the mineral because it changes the hydrophobicity and the hindering effect on collector absorption. When a galvanic contact between sulfide and iron, “ the values of mixed potentials were in the vicinity of xanthate/dixanthogen redox potential …, when in contact with iron, the mixed potentials shifted towards a more negative value, making them too reducing for xanthate collector to be absorbed effectively” (1).

Studies have shown that galvanic interactions between grinding media and minerals in aqueous solutions are electrochemical in nature. In other words, “galvanic interactions occur in every grinding media-sulfide mineral system” (1). Reducing this interaction has been proposed by substituting forged steel grinding media with inert grinding media into the system. From a cost-effective approach, ceramic and stainless- steel grinding media are not suitable to replace forged steel grinding media. On the other hand, high chrome white iron initially “corrodes to form a very hard chrome oxide layer which effectively then prevents the exchange of electrons and halts the corrosion process” (1). The interaction of chrome frees oxygen in the grinding media – sulfide mineral system oxidiing sulfide minerals such as pyrite and pyrrhotite and, hence, inhibiting the flotation of these minerals.

Specifications of Experiment

1. Comminution of Sample, with Screen Analysis

All the material received was initially crushed through a BICO INC jaw crusher, followed by a BRAUN chipmunk VD67 crusher, and finally passed through a BRAUN roll crusher. The effect of this crushing circuit produced a direct reduction of as-received material to 90% passing a -10-mesh screen in stages. The nominal sies were as follow:

3 1 ⁄8 "(9.5 mm) from the jaw crusher, ⁄4 " (6.7 mm) from the chipmunk crusher and 10 mesh (1.7 mm) from the roll crusher. After the last crushing stage, the material was split by using the sample technique called splitting. Pouring material through a splitter, the material was divided into two equal parts. One of the samples was accepted as a

20 representative sample of the ore while the other one was rejected. The ore was split four times until a representative sample of 7.48 kg was obtained for subsequent experiments.

2. Milling Sie

500-g sample of -10 mesh was placed in a laboratory ball mill with a “seasoned charge” (a ball mill charged with a distribution of ball mill sies to mimic steady sate ball charge composition). 500-ml of water were added, and the ball mill was sealed and placed on the rolling table for 30-minutes, 45-minutes and 60-minutes. After each run, the material was dried, and screened to establish the 80% passing sie after milling.

These tests were repeated as necessary to determine the milling time, and effectively achieve a P80 sie of 100 mesh. It was anticipated to conduct 5 to 8 grinding tests and estimate the optimal grinding time. This testing section required a total sample weight between 2.5 to 4.0 kg. To determine the P80, a 100-mesh screen which is equal to 149 microns or 0.149 mm was used to determine this value. A technique known as wet screening was utilied to obtain the +100-mesh percentage left on the screen. After washing the material, a rough visual observation was used to determine if a P80 could be achieve using a 100-mesh screen. At 45 minutes, the result showed that the material was sufficiently liberated to apply rougher flotation and separate valuable material from gangue material through surface chemistry modifications.

3. Flotation Experiment

500-g samples of the -10-mesh material was milled with 500-ml of water for 45 minutes, as determined in step 2. During the milling step, half of the reagents were added to have a thorough effect to recover material. The slurry was poured from the mill and

21 placed in a 1-L Denver flotation cell with an 11-1200 rpm range. Additional reagents including frothers were added, and the pulp was conditioned for 10 minutes prior to air injection. After the air was injected, flotation began and continued for approximately 15 minutes by collecting the froth in a small pan using a spatula. The actual flotation time varied depending on the observed kinetics of flotation of the material and was dependent upon when mineral recovery was complete.

Samples from concentrate and tail were filtered, dried, and split for XRD analysis. A total of 10 flotation tests were conducted simulating each reagent schemes discussed in this section. Running this experiment required an approximate consumption of 10 lb. of material from the collected sample.

Results and Discussion

Early experiments were conducted to determine the grinding time that would liberate the most valuable material from the gangue material or waste. After pulveriing the crushed material to achieve -10 mesh nominal sie, the material was ground for a certain amount of minutes (30, 45 and 60). Choosing these time intervals was based on ore characteristics such as hardness, and the mineralogical composition of the ore which included molybdenite, chalcopyrite, pyrite, and quart as the main silicate mineral. To evaluate what grinding time would give a good mineral liberation, a 100-mesh screen was used to separate coarse particles from fine particles. After sieving the sample for each time interval, the retaining sample on the screen was weighed. Thus, 45 minutes of grinding would achieve an 80 % passing or P80 sie of approximately 100 mesh based on visual observation from the material left as the coarser sie on the sieve.

During the grinding stage, reagents were added following the different reagent schemes selected to evaluate the chalcopyrite recovery and pyrite rejection. The addition of reagents in the grinding circuit was a representation of how mill plants carry out their processes as well. It gives more time for reagents to act with the minerals prior to the flotation stage and allows for reaction with fresh mineral surfaces. Table 3 shows a summary of the reagents used for the grinding and flotation stages. As a rule of thumb, half of the dosages were introduced prior to the grinding stage and the other half during the flotation stage. After ten minutes of conditioning, air was injected to the system to create bubbles collecting the concentrate known as rougher concentrate.

Table 3: Reagent schemes used to improve copper recovery and depress pyrite (*).

Reagent Standard Scheme #1 Scheme #2 Scheme #3 Scheme #4

CuFeS2 collector #1 Xanthate AERO 3302 Dithiophosphate AERO 3302 Thiocarbamate

CuFeS2 collector #2 ~ AERO3302 AERO 3302

MoS2 collector Fuel Oil Fuel Oil Fuel Oil Fuel Oil Fuel Oil

FeS2 depressant #1 Ca(OH)2 Na2SO3 K2Cr2O7 ZnSO4 Na2SO3

FeS2 depressant #2

SiO2 depressant Frother MIBC MIBC MIBC MIBC MIBC

Reagent Scheme #4.1 Scheme #4.2 Scheme #5 Scheme #6 Scheme #7

CuFeS2 collector #1 Thiocarbamate Thiocarbamate AERO 3302 AERO 3302 AERO 3302

CuFeS2 collector #2 AERO 3302 AERO 3302

MoS2 collector Fuel Oil Fuel Oil Fuel Oil Fuel Oil

FeS2 depressant #1 K2Cr2O7 K2Cr2O7 K2Cr2O7 Ca(OH)2 K2Cr2O7

FeS2 depressant #2 Ca(OH)2 Ca(OH)2

SiO2 depressant Na2SiO3 Frother MIBC MIBC MIBC MIBC MIBC (*) Refer to Table 1 for dosages used to recover chalcopyrite and reject pyrite.

An ore sample of 500 grams was sent for an assay analysis using inductively

coupled plasma (ICP) spectroscopy to detect and measure the elements within the ore

sample. Results were reported in percentages corresponding to each element found in the

mineral sample (See Table 4). These results were used to calculate the mineral percentage

for the ore sample prior to reagent addition in the grinding and flotation stages. To

accurately calculate FeS2 present in the ore, iron within the chalcopyrite was subtracted

from the overall pyrite in the ore, and it was tabulated in table 4.

Table 4: Assay analysis of a 500-g after using ICP analytical method.

Metal Grade Mineral Grade Cu 0.036% CuFeS2 0.104% Mo 0.023% MoS2 0.038% Fe 1.74% FeS2 3.67%

Flotation tests were performed under the same ambient conditions of temperature, pressure, mass, volume, and equipment. All tests were conducted with the same basic procedure, with a 10-minute conditioning time to ensure consistency of the tests. A high pH is an important factor in the depression of pyrite in many flotation circuits; one test was conducted to establish a standard pyrite depression so that all other experiments could be compared with it for pyrite rejection (See Fig 7 and 8).

Qualitative information was gathered by collecting a rougher concentrate from a

1-L Denver flotation cell. The collected samples for each scheme of reagents were analyed implementing x-ray diffraction analysis (XRD) and the results were tabulated in

Table 5.

Table 5: Percentage for each mineral found in the rougher concentrate corresponding to each scheme.

Auto-qualitative Semi-qualitative Scheme CuFeS2 MoS2 FeS2 CuFeS2 MoS2 FeS2 Standard 0.60% 5.6% 9.1% 0.4% 28.2% 6.9% Scheme 1 2.80% 10.3% 58.8% 2.6% 16.1% 55.1% Scheme 2 4.2% 12.4% 41.5% 2.8% 21.3% 27.1% Scheme 3 3.6% 19.1% 48.4% 3.0% 32.3% 40.5% Scheme 4 14.7% 8.4% 46.1% 14.0% 12.8% 43.9% Scheme 4.1 4.6% 31.1% 20.1% 3.8% 43.5% 16.5% Scheme 4.2 2.20% 9.0% 53.3% 2.00% 14.6% 50.0% Scheme 5 1.5% 16.8% 42.7% 1.4% 25.6% 38.2% Scheme 6 2.5% 8.3% 18.8% 0.9% 3.2% 7.2% Scheme 7 6.1% 11.8% 63.1% 4.1% 40.6% 42.6%

The XRD software allows to conduct two side-by-side analyses of every sample.

After the initial scan was performed, the minerals are identified by matching with standard scans in the software data. The auto qualitative (AQ) analysis averages all the peaks and normalies them to add to 100%. However, with the semi qualitative (SQ) analysis, one of the peaks was chosen as the standard which all others are normalied to match. The percentages sum up to 100%, but there was a small weighing factor associated with the SQ analysis. The decision to use both auto and semi qualitive data was to ensure that the recovery for copper and iron were correct, but most importantly, trends (not quantitative numbers) were the main information used for the comparison of reagent schemes. When XRD was set in auto-qualitative mode, the highest peak of reference was quart due to its abundance within the concentrate and all other minerals were factored compared with it. Yet, in semi-qualitive mode, the analysis was set to

26 consider molybdenite as the reference point giving different percentages for each mineral.

To obtain a thorough qualitive analysis, both XRD settings were included in this result

(See Table 3 and Appendix).

The results from Table 5 were used to calculate the metal percentage and thus estimate the metal recovery for each scheme of reagents (See Table 6). As a result, the data was tabulated to observe and determine what combination of reagents might provide the desired results.

Table 6: Recovery for copper and iron in the rougher concentrate using auto and semi qualitative settings from XRD analysis.

Recovery AQ Recovery SQ Scheme Cu Fe Cu Fe Standard 12.00% 5.28% 8.00% 3.98% Scheme #1 52.25% 31.43% 48.51% 29.44% Scheme #2 70.29% 20.57% 46.86% 13.45% Scheme #3 76.87% 30.11% 64.06% 25.19% Scheme #4 76.35% 31.86% 72.71% 30.34%

Scheme #4.1 12.21% 1.70% 10.09% 1.40% Scheme #4.2 21.58% 14.92% 19.62% 13.98% Scheme #5 36.07% 29.18% 33.66% 26.13%

Scheme #6 78.78% 13.65% 28.07% 19.65% Scheme #7 86.83% 26.53% 58.36% 17.91%

Commonly used in concentrators dedicated to extract copper concentrates, xanthate and calcium sources are the best option to recover chalcopyrite and reject pyrite.

As part of this thesis work, limestone and xanthate were used to float and collect the rougher concentrate. The collected concentrate was further analyed through XRD to obtain the mineral percentages (See Fig. 5 and 6). Those percentages served to calculate the metal recovery for the standard scheme (See Table 5 and 6). Studies have shown that

27 this scheme of reagents enhances copper recovery and iron rejection; yet low recovery was observed (see Fig 7 and 8). It can be speculated that other factors may be influencing the recovery impacting the selectivity of this scheme resulting in a low recovery (See Fig

7 and 8, also refer to Appendix). Moreover, it has been documented that the milling environment influences the galvanic interactions in the grinding media – sulfide mineral system. During grinding stages, grinding media, ore and reagents are interacting with each other which might cause a coating effect of chalcopyrite surfaces with iron ions and pyrite surfaces with copper (I). For instance, metallurgical improvements at Kagara’s

Mount Garnet mine found this galvanic interaction within the grinding media – sulfide mineral system. It was documented that a hindering effect occurs because “ the values of mixed potentials were in the vicinity of xanthate/dixanthogen redox potential …, when in contact with iron, the mixed potentials shifted towards a more negative value, making them too reducing for xanthate collector to be absorbed effectively” (1). Nonetheless, further studies including oxidation potentials of the xanthate/dixanthogen redox potential are needed to understand the interaction of a grinding media – sulfide mineral system.

This measurement can be done using an oxidation-reduction potential sensor (ORP), and it can indicate the amount of oxygen present in the mixture by reading the voltage in milliVolts (mV).

Figure 5: Standard scheme modifying pH by adding lime to depress pyrite (Auto-qualitative).

Figure 6: Standard scheme modifying pH by adding lime to depress pyrite (Semi-qualitative).

The results obtained from XRD analyses using auto and semi qualitive settings were tabulated to calculate the metal recovery based on the percentages of mineral found on the rougher concentrate. Table 5 contains the metal recovery for copper and iron from

29 auto and semi qualitive setting. To observe the recovery of copper compared to iron, two comparison trends or bar graphs were constructed (Fig 7 and 8). As it has been

- mentioned, hydrated lime (Ca(OH)2) with xanthate (ROCSS 2M) as the collector are the most common reagents for the depression of pyrite (main source of iron) and, to collect chalcopyrite (main source of copper), respectively. In this experiment, xanthate with limestone represent the standard scheme when compared to the other schemes in study

(See Fig 7 and 8). pH values for the standard scheme were reported as 8.8 prior to lime addition and 10.5 during the flotation portion of the experiment. On the other hand, the other schemes were conducted with pH values ranging between 8.45-10.5 to observe the versatility of AERO 3302 promoter.

Based on the low recoveries observed in the standard lime/xanthate test, it would be unfair to say the mix of xanthate as copper collector and limestone as pH modifier to depress pyrite does not work, yet it does suggest that iron oxidation from the grinding may have influenced the copper recovery. For this reason, a grinding media affect was considered since it naturally occurs in aqueous solutions when grinding media, ore and reagents interact with each other. Bruckard and Woodcock considered that corrosive wear of grinding media increases due to electrochemical interactions, but it also changes the surface properties and pulp chemistry of the grinded ore affecting the subsequent flotation (2). In their studies along with those from The Australasian Institute of Mining and Metallurgy, both suggested that replacing forged steel grinding media with inert grinding media would yield better copper recovery and iron rejection. High white chrome grinding media was found to reduce the galvanic interaction among the grinding media – sulfide mineral system, and one of the objectives was to implement the use of grinding

30 media made of chrome. Yet, manufacturing constraints would not allow it since chrome grinding media was expensive to produce for small scale testing facilities. Hence, tests were conducted using potassium dichromate (K2Cr2O7) (a chemical mimicking the effect of chrome-iron media) as a depressant instead of replacing the forged steel grinding media was considered. Previously stated, the grinding environment enhances the galvanic interactions among sulfide minerals and grinding media, thus the anodic contribution of

K2Cr2O7 was crucial and important to see its effect on the depression of pyrite. Along

- with this modification, xanthate (ROCSS 2M) was replaced for a more selective reagent favoring the recovery of molybdenite. A commercial reagent manufactured by Cytec

Industries Inc. called AERO 3302 promoter was implemented to carry on the subsequent flotation tests.

Based on the results shown in Fig. 7 and 8, schemes 1, 3 and 4 show a higher iron recovery in comparison to scheme 2, 4.1, 4.2, 5, and 7. This difference in iron recovery was due to the addition of potassium dichromate (K2Cr2O7) into those schemes. Overall recovery of iron in auto and semi qualitative settings from XRD analyses shows percentages approximately below 30%. Nonetheless, schemes 2, 4.1, 4.2, 5 and 7 presents an iron recovery slightly lower when compared to schemes 1, 3, 4, 6 and the standard scheme. For trending purposes, schemes 2, 4.1, 4.2, 5 and 7 demonstrate that potassium dichromate acts as a depressant when added to the milling system (ball mills).

It has been demonstrated that high chrome grinding media initially “corrodes to form a very hard chrome oxide layer which effectively then prevents the exchange of electrons and halts the corrosion process” (1). Due to this interaction, more oxygen is available into the system favoring the oxidation of sulfide minerals such as pyrite and may contribute to

31 preventing its flotation. Notwithstanding, high chrome grinding media was not used, potassium dichromate exhibits a depressing effect by actuating as the counter part of the galvanic interaction between dissolved iron ions and the dichromate ion. This reaction allows the system to have more oxygens to oxidie the surface of pyrite and, hence, its further depression due to pyrite’s affinity for water. Additionally, the availability of oxygen due to the use of chrome as grinding media or depressant (the case of this research) helps to assist in the effectiveness of the collector absorption and mineral floatability.

% Recovery: Auto-Qualitative

100% 90%

80% 70%

60% 50%

40% 30%

20% 10%

Cu Fe

Figure 7: Comparison of Cu and Fe recovery using different schemes of reagents. In this figure, the recovery is from an auto-qualitative perspective set by XRD parameters to obtain percentage recovery for each mineral containing Cu and Fe.

% Recovery: Semi-Qualitative 100% 90% 80% 70% 60% 50% 40% 30% 20% 10% 0%

Cu Fe

Figure 8: Comparison of Cu and Fe recovery using different schemes of reagents. In this figure, the recovery is from a semi-qualitative perspective selecting molybdenum as the reference point to obtain percentage recovery for each mineral containing Cu and Fe.

Scheme 6 replicates the standard scheme but using AERO 3302 as the copper collector. This reagent is known for being insoluble due to its oily characteristic.

Standing by itself and compared to the standard scheme, AERO 3302 has a higher selectivity of copper particles. It also indicates that because of its oily characteristic causes a coating effect during the grinding stage. This effect can help to depress pyrite and have more copper in the rougher concentrate, because it was not as reactive with the iron surface as copper surfaces. Likewise, all schemes (see Table 1) except the standard scheme have AERO 3302 as copper collector along with other collectors such as thiocarbamate and dithiophosphate. Comparing the rougher concentrates for each scheme, AERO 3302 has a better selectivity compared to xanthate. Xanthate presents a disadvantage requiring a pH higher than 9 to depress pyrite and yield better results for chalcopyrite recovery. On the other hand, AERO 3302 showcases a more versatile

33 approach since it can work in a wider range of pH. Controlling pH was not a major factor in this experiment, it can be observed that AERO 3302 floated more copper into the rougher concentrate.

To better understand the results from figures 7 and 8, table 6 was condensed into a selectivity ratio between metal recoveries (see table 7). This selectivity ratio was based on the copper/iron metal recovery from each XRD setting, and it helped to determine what schemes were having a better recovery when compared to the standard scheme.

Table 7: Selectivity ratio for metal recoveries.

Scheme SQ Ratio AQ Ratio Standard 2.0 2.3 Scheme #1 1.6 1.7 Scheme #2 3.5 3.4 Scheme #3 2.5 2.6 Scheme #4 9.5 9.5 Scheme #4.1 7.2 7.2 Scheme #4.2 1.4 1.4 Scheme #5 1.3 1.2 Scheme #6 1.4 5.8 Scheme #7 3.3 3.3

It was noticed that schemes 1, 4.2, 5 and 6 were being less selective than the standard scheme. The metal recoveries were relatively close indicating that minerals (CuFeS2 and

FeS2) containing these elements were proportionally recovered in the concentrate. On the other hand, schemes 2, 3, 4, 4.1 and 7 showed to be more selective than the standard test.

These schemes showed to be more selective for chalcopyrite and a higher rejection for pyrite in the rougher concentrate. This ratio served to visualie the effectiveness of each scheme with respect to each other and the standard scheme (see Figure 9).

Selectivity Ratio: Metal Recovery 10.0 9.0 8.0 7.0 6.0 5.0 4.0 3.0 2.0 1.0 0.0

Metal Recovery AQ Metal Recovery SQ

Figure 9: Selectivity ratio summariing metal recovery for each scheme.

The XRD analyses and metal recovery comparison for each scheme suggested that using 3302 AERO promoter at approximately 4.50 – 9.00 lb./ton and dichromate as a depressant at approximately 0.015 lb./ton improve the copper recovery while reducing the amount of iron present in the rougher concentrate. Along with that, the combination of different reagents might have contributed to improved recoveries. Notwithstanding, further analyses focusing on the interaction of reagent-reagent, reagent-ore and reagent- grinding media must be considered as part of the study.

Further Studies

To understand galvanic interactions between grinding media – sulfide mineral system, oxidation tests must be included. This can help to understand the different redox potentials occurring prior to and during the addition of reagents to the system.

Additionally, it would give an insight of how the anodic contribution reacts with the

35 grinding media instead of its replacement. Another aspect that can be further analyed is the interaction between limestone, dichromate and grinding media in a system and how it affects the depression of pyrite when subsequent flotation is applied.

Conclusions

The following conclusions may be drawn based on the results of this study:

1. Using potassium dichromate as a depressant instead of replacing the forged steel

grinding media shows that pyrite was depressed resulting in more copper at the

rougher concentrate.

-2 2. The anodic contribution of potassium dichromate (Cr2O7 ) seems to interact with

the dissolved iron ions in the system freeing more oxygen to oxidie pyrite.

Further studies including redox potential should be included to observe this

possible phenomenon.

3. AERO 3302 promoter has a stronger selectivity compared to xanthate due to its

wide range of pH. Xanthate needs a certain pH condition (above 9) to be more

selective.

4. The combination of AERO 3302 and potassium dichromate seemed to produce

better recoveries for copper and low iron without needing to control pH through

the addition of lime.

5. The grinding environment might have a critical role when it comes to the galvanic

interaction among sulfide minerals and grinding media, as shown by the pyrite

rejection when dichromate salts were added to the flotation tests.

6. As stated by researchers, high chrome grinding media due to its inert nature has

shown to favor the availability of oxygen in the system which further in the

reaction oxidies pyrite causing its depression. Potassium dichromate as

37 depressant might have similar effects allowing more oxygen to further oxidie pyrite during the rougher flotation stage.

Reference

1. AusIMM, T. (2006). Metallurgical improvements at kagara’s mount garnet mine

through the use of high chrome grinding media. (pp.1-1) The Australasian Institute of

Mining and Metallurgy (The AusIMM).

2. Bruckard, W. J., Sparrow, G. J., & Woodcock, J.T. (2011). A review of the effects of

the grinding environment on the flotation of copper sulphides. International Journal of

Mineral Processing, 100(1), 1-13.

3. Moslemi, H., & Gharabaghi, M. (2017). A review on electrochemical behavior of pyrite

in the froth flotation process. Journal of Industrial and Engineering Chemistry, 47, 1-

18.

4. Liu, G., Lu, Y., Zhong, H., Cao, Z., & Xu, Z. (2012). A novel approach for preferential

flotation recovery of molybdenite from a porphyry copper-molybdenum ore. Minerals

Engineering, 36, 37-44.

5. Davenport, W. G., & Biswas, A. K. 1. (2002). Extractive metallurgy of copper (Fourth

ed.). Oxford:Davers, Mass;: Pergamon.

6. Huang, K. (2015). Validation and application of a first principle flotation model.

Virginia Polytechnic Institute and State University, Blacksburg, Virginia.

7. Cytec Industries Inc. (2002). Mining chemicals handbook (Revis ed). Bound Brook,

N.J.: Cytec Industries Inc.

8. Mu, Y., Peng. Y. & Lauten, R. A., (2016). Depression of pyrite in the flotation of

chalcopyrite using biopolymer depressants. The Canadian Institute of Mining,

Metallurgy and Petroleum, 1-11.

9. RABATHO, J., TONGAMP, W., KATO. J., HAGA, K., TAKASAKI, Y., &

SHIBAYAMA, A. (2011). Effect of flotation reagents for upgrading and recovery of

cu and mo form mine tailing by flotation. Resources Processing, 58(1), 14-21.

10. Guo, B., Pen, Y., & Parker. G. (2016). Electrochemical and spectroscopic studies of

pyrite—cyanide interactions in relation to the depression of pyrite flotation. Minerals

Engineering, 92, 78-85.

11. Mu, Y., Peng, Y., & Lauten, R. A. (2016). The depression of pyrite in selective flotation

by different reagents systems – A literature review. Minerals Engineering, 96-97, 143-

156.

12. Fornasiero, D., & He, S. H. (n.d). Minimising copper activation of iron sulfide minerals

using grinding. Ian Wark Research Institute, University of South Australia, 1048-1053.

13. Cullity, B. D. (1978). Elements of x-ray diffraction (Second ed.). Reading, Mass:

Addison-Wesley Pub. Co.

Appendix

Figure 10:Standard scheme using xanthate, fuel oil and limestone as flotation reagents (AQ).

Figure 11: Standard scheme using xanthate, fuel oil and limestone as flotation reagents (SQ).

Figure 12: Scheme 1 using thiocarbamate, AERO 3302, fuel oil and sodium sulfite as flotation reagents (AQ).

Figure 13: Scheme 1 using thiocarbamate, AERO 3302, fuel oil and sodium sulfite as flotation reagents (SQ).

Figure 14: Scheme 2 using AERO 3302, fuel oil and potassium dicromate as flotation reagents (AQ).

Figure 15: Scheme 2 using AERO 3302, fuel oil and potassium dichromate as flotation reagents (SQ).

Figure 16: Scheme 3 using AERO 3302, fuel oil and inc sulfate as flotation reagents (AQ).

Figure 17: Scheme 3 using AERO 3302, fuel oil and inc sulfate as flotation reagents (SQ).

Figure 18: Scheme 4 using thiocarbamate, AERO 3302, fuel oil and sodium sulfite (AQ).

Figure 19: Scheme 4 using thiocarbamate, AERO 3302, fuel oil and sodium sulfite as flotation reagents (SQ).

Figure 20: Scheme 4.1 using thiocarbamate, AERO 3302, fuel oild and potassium dichromate as flotation reagents (AQ).

Figure 21: Scheme 4.1 using thiocarbamate, AERO 3302, fuel oil and potassium dichromate as flotation reagents (SQ).

Figure 22: Scheme 4.2 using thiocarbamate, AERO 3302, fuel oil and potassium dichromate as flotation reagents (AQ).

Figure 23: Scheme 4.2 using thiocarbamate, AERO 302, fuel oil and potassium dichromate as flotation reagents (SQ).

Figure 24: Scheme 5 using AERO 3302, fuel oil, potassium dichromate and lime as flotation reagents (AQ).

Figure 25: Scheme 5 using AERO 3302, fuel oil, potassium dichromate and lime as flotation reagents (SQ).

Figure 26: Scheme 6 using AERO 3302 and lime as flotation reagents (AQ).

Figure 27: Scheme 6 using AERO 3302 and lime as flotation reagents (SQ).

Figure 28: Scheme 7 using AERO 3302, fuel oil, potassium dichromate, lime and sodium sulfite as flotation reagents (AQ).

Figure 29: Scheme 7 using AERO 3302, fuel oil, potassium dichromate, lime, and sodium sulfite as flotation reagents (SQ).