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Chloride leaching for chalcopyrite Liddicoat, Jenni Anne 2003

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Chloride Leaching for Chalcopyrite by Jenni Anne Liddicoat B.Sc., The University of Sydney, 1997 A THESIS SUBMITTED IN PARTIAL F U L F I L M E N T OF THE REQUIREMENTS FOR THE DEGREE OF MASTER OF SCIENCE in THE F A C U L T Y OF G R A D U A T E STUDIES Department of Metals and Materials Engineering We accept this thesis as conforming to the required standard The University of British Columbia December 2003 © Jenni Liddicoat, 2003 In p r e s e n t i n g t h i s t h e s i s i n p a r t i a l f u l f i l m e n t o f t h e r e q u i r e m e n t s f o r an a d v a n c e d d e g r e e at t h e U n i v e r s i t y o f B r i t i s h C o l u m b i a , I ag ree t h a t t h e L i b r a r y s h a l l make i t f r e e l y a v a i l a b l e f o r r e f e r e n c e and s t u d y . I f u r t h e r a g r e e t h a t p e r m i s s i o n f o r e x t e n s i v e c o p y i n g o f t h i s t h e s i s f o r s c h o l a r l y p u r p o s e s may be g r a n t e d by t h e head o f my depar tment o r by h i s o r her r e p r e s e n t a t i v e s . I t i s u n d e r s t o o d t h a t c o p y i n g or p u b l i c a t i o n o f t h i s t h e s i s f o r f i n a n c i a l g a i n s h a l l not be a l l o w e d w i t h o u t my w r i t t e n p e r m i s s i o n . Department o f The U n i v e r s i t y o f B r i t i s h C o l u m b i a V a n c o u v e r , Canada A B S T R A C T Two new process flowsheets have been developed which combine chloride leaching for chalcopyrite with solvent extraction, to selectively transfer copper to a conventional sulfate electrowinning circuit. Both models were designed to address the common chloride hydromet problem of product impurity, and they differ with respect to iron deportment. Chloride leaching offers significant advantages for copper hydrometallurgy including increased solubility, increased rates of leaching and stabilization of Cu(I) through chloride complexation. A mass balance was completed for both models and enabled the determination of recycle leach liquor concentrations of copper and iron species for the countercurrent leach tests. To assist with solid-liquid (S-L) separation a bench size mini-thickener was designed and tested in the continuous countercurrent,leach experiments. One mini-thickener was used for leach stage 1 (LSI) S-L separation in a smooth, clean and semi-continuous manner. The leach tests were performed using a chalcopyrite concentrate, from Antamina in northern Peru, which contained a low to moderate amount of gangue material. The successful completion of mass balances for both circuits demonstrates potential for these two new chloride leach flowsheets particularly in addressing current purification difficulties and waste management issues. The goethite model leach experiments, in which oxygen was injected to leach stage 2 (LS2) to aid iron oxidation and precipitation, were unsuccessful in achieving >95% copper extraction in the atmospheric conditions tested. It is believed that further intensification of conditions could produce >95% copper extraction. The hematite model, in which no air is added to the leach, was successful in demonstrating the effect of varying particle size, leach time and temperature to achieve copper extractions >95%. Three hematite process variations achieved >95% copper extraction and these involved fine grinding of the concentrate, and either a 3h residence ii time at 95C (98% extraction), a 2h residence time at 95C (96% extraction) or a 3h residence time at 85C (96% extraction). One final experiment, based on the hematite model, was successfully conducted at optimum leach conditions employing Rosario concentrate from Chile, which contains chalcopyrite, chalcocite and a significant amount of pyrite. Copper extraction exceeded 99% for this experiment. i i i TABLE OF CONTENTS A B S T R A C T i i T A B L E O F C O N T E N T S iv L I S T O F F I G U R E S v i i L I S T O F T A B L E S ix A C K N O W L D E G E M E N T S x i 1 I N T R O D U C T I O N 1 2 L I T E R A T U R E R E V I E W 4 2.1 Aqueous copper and chloride chemistry 4 2.1.1 Chloride complexes 4 2.1.2 Solubility 8 2.1.3. Ionic Act ivi ty 10 2.1.4 Redox Chemistry 12 2.2 Chalcopyrite leaching in chloride 17 2.2.1 Thermodynamics and kinetics 17 2.2.2 Sulfur 23 2.2.3. Iron Deportment 25 2.3 Overview o f chloride leach processes for copper 29 2.4 Summary 34 3 E X P E R I M E N T A L 35 3.1 Mass balance spreadsheet preparation for the leach models 35 3.2 Feed concentrates 35 3.3 Feed solutions 37 3.3.1 Feed liquor for the hematite model 38 3.3.2 Feed liquor for the goethite model 38 3.3.3 Feed liquor analysis 38 3.4 Experimental Apparatus 39 3.4.1 Leach vessels 39 3.4.2 Temperature control 40 3.4.3 Solid-liquid separation 40 3.4.4 Compressed air/oxygen addition to L S 2 43 iv 3.5 Experimental procedures 44 3.6 p H and redox potential measurements 48 3.6.1 p H 48 3.6.2 Redox potential 49 3.7 Sampling and Analysis 49 3.7.1 Sample Collection 49 3.7.2 Preparation and Analysis 50 3.7.3 Calculation of copper and iron extraction 51 4 R E S U L T S A N D D I S C U S S I O N 52 4.1 Excel mass balance outputs for the hematite and goethite process flowsheets 52 4.1.1 The hematite process 53 4.1.2 The goethite process 55 4.1.3 Feed liquor outputs 57 4.2 Overview of continuous countercurrent leach experiments 58 4.3 Results of hematite model experiments 62 4.3.1 Potential, p H and free acid 63 4.3.2 Copper and iron extraction 66 4.3.3 Speciation 69 4.3.4 Hematite experiment discussion 71 4.4 Results of goethite model experiments 71 4.4.1 Potential, p H and free acid 73 4.4.3 Speciation 78 4.4.4 Goethite experiment discussion 79 4.5 Application of hematite leach model to Rosario concentrate 79 4.5.1. Potential, p H and free acid 80 4.5.2 Copper and iron extraction 81 4.6 Sulfur balance 83 4.7 Trace Metals 85 5 C O N C L U S I O N S A N D R E C O M M E N D A T I O N S 86 5.1 Mode l mass balances 86 5.2 Implementation of the mini-thickener 86 5.3 Countercurrent leach experiments 87 5.4 Recommendations for further work 87 6. R E F E R E N C E S 89 Sample identification guide for Appendices 1 and 2 93 Appendix 1 Certificates of Analysis (IPL) 94 Appendix 2 Certificates of Analysis (ChemMet) 121 Appendix 3 Certificates of Analysis (Malvern Mastersizer) 131 Appendix 4 X R D Spectra 135 Appendix 5 Methods 136 Appendix 6 Experimental Logs 143 Appendix 7 Calculations 147 Appendix 8 Mass balances for experimental accountability 151 Appendix 9 Experimental Profiles 164 vi L I S T O F F I G U R E S Figure 1 The Hematite Process 3 Figure 2 The Goethite Process 3 Figure 3 C u - H 2 0 E h p H Diagram at 25°C (1 atm, 1M) [12] 6 Figure 4 C u - C l - H 2 0 E h p H Diagram at 25°C (1 atm, 1M) [12] 6 Figure 5 C u - H 2 0 E h p H Diagram at 25°C (1 atm, C u 1 M , C l 5M)[12] 6 Figure 6 C u C l Solubility [1] 9 Figure 7 PbC12 Solubility [1] 9 Figure 8 Altered redox potentials in chloride solution [3] 13 Figure 9 E h vs L o g aCl [23] 15 Figure 10 Cu-Fe-S-H20 Eh-pH diagram at 25°C [24] 18 Figure 11 Cu-Fe-S-Cl -H20 Eh-pH diagram at 25°C [24] 19 Figure 12 Chloride leach processes for copper 30 Figure 13 Experimental Set-up 39 Figure 14 Mini-thickener design 42 Figure 15 Photo of mini-thickener 42 Figure 16 Photo of titanium rake 42 Figure 17 Photo of countercurrent leach set-up 44 Figure 18 Mass and volume transfers 46 Figure 19 Thickening operation 47 Figure 20 Hematite Process 53 Figure 21 Goethite Process 55 Figure 22 E profile for H4(75um, 3h, 95C) 64 Figure 23 E profile for H5(41um, 3h, 95C) 64 Figure 24 p H profile for H4(75um, 3h, 95C) 64 Figure 25 p H profile for H5(41um, 3h, 95C) 64 Figure 26 H4: Cu & Fe (%) Extraction ; 68 Figure 27 H5 : C u and Fe (%) Extraction 68 Figure 28 H4: Concentration vs Time 68 Figure 29 H5 : Concentration vs Time 68 Figure 30 H4: C u and Fe in LS2 Residue 68 vii Figure 31 H5 : C u and Fe in L S 2 Residue 68 Figure 32 E profile for G2(75um, 2h, 85) 74 Figure 33 E profile for G3(41um, 3h, 95C) 74 Figure 34 p H profile for G2(75um, 2h, 95C) 74 Figure 35 p H profile for G3(41um, 3h, 95C) 75 Figure 36 G l : C u & Fe (%) Extraction 77 Figure 37 G2: C u & Fe (%) Extraction 77 Figure 38 G3: C u & Fe (%) Extraction 78 Figure 39 E profile for H9(37um,3h,95C) 81 Figure 40 p H profile for H9(37uni,3h,95C) 81 Figure 41 C u & Fe (%) Extraction for H9 82 Vlll LIST OF TABLES Table 1 Metal-chloro complexes with increasing chloride concentration [3] 7 Table 2 Mean activity (a) of H + and Cl" in 2 M HC1 and several salts [1] 11 Table 3 Summary of chloride leaching kinetics 22 Table 4 Summary of iron compound precipitation [36,37] 28 Table 5 Process Review: Chloride Leach Summary Table 31 Table 6 Process Review: Summary Table 32 Table 7 Cu , Fe and S % in feed concentrates 36 Table 8 Feed concentrate particle size 37 Table 9 Hematite Process: L S 2 streams 53 Table 10 Hematite Process: L S I streams 54 Table 11 Hematite Process: Solvent Extraction streams 54 Table 12 Hematite Process: Autoclave streams 55 Table 13 Goethite Process: LS2 streams 56 Table 14 Goethite Process: L S I streams 57 Table 15 Goethite Process: Solvent extraction streams 57 Table 16 Feed liquor concentrations 57 Table 17 Order of experiments 59 Table 18 Hematite model - feed conditions 62 Table 19 E , p H and free acid output for hematite experiments 65 Table 20 Results of hematite model experiments 67 Table 21 Total reduced species of exit P L S and L S 2 liquors 69 Table 22 Feed conditions for goethite model experiments 72 Table 23 E , p H and free acid output for goethite experiments 74 Table 24 Output concentrations (solution and solids) for goethite model experiments... 76 Table 25 Total reduced species in goethite experiments 79 Table 26 Feed conditions for Rosario experiment 80 Table 27 E , p H and free acid 81 Table 28 Output concentrations (solution and solids) for the Rosario experiment 82 Table 29 Sulfur species in Final LS2 residues 83 Table 30 Sulfur recovery 84 ix Table 31 Trace metal analysis 85 Table 32 Sample identification guide 93 Table 33 Concentrate sample identification 93 Table 34 Comparing IPL free acid results with test titration value 138 x ACKNOWLDEGEMENTS I would like to sincerely thank Professor David Dreisinger for both the opportunity to undertake this masters and for the guidance over the past two years. The Tim-Tam biscuits brought from Australia were also very much appreciated. I also thank Falconbridge Ltd for the opportunity and importantly for the project funding. In particular Mohammed Buarzaiga, from Falconbridge, was a gem in assisting with the design and prompt manufacture of the mini-thickener and rake component. Other department faculty and staff have also been helpful in advising and assisting with not only my studies but also my social enjoyment. I owe thanks to David Dixon for the chemistry instruction, Steve Cockcroft (in particular for the B B Q and Softball funding), and the kind staff in the M M A T workshop and stores who were always there to help me put things back together and working again. These guys deserve a special mention Serge, Ross, Carl, Dave and Glenn. I would also like to thank the U B C Department of Min ing and Department of Geological Sciences for their cooperation in allowing me to use their analytical instruments. I owe a great deal of thanks to Prabhjit Bhatia, my trusty research aid, who took care of my experiment during the daytime shift, when I was sleeping to rejuvenate for the night shift. The Department of Metals and Materials Engineering at U B C has been an enjoyable place to work, and the other graduate students and staff have indeed made it a friendly place to be. I would like to say a special thanks to all of my lab and office mates in the hydromet group, both passing and permanent. In particular, there was one thoughtful visitor who kept me going during my early struggles with apparatus and procedure development, and who later helped me to revise my thesis. x i 1 INTRODUCTION Within the past 150 years a number of hydro-metallurgical techniques have been proposed for the extraction of base metals, copper (Cu), nickel (Ni), lead (Pb) and zinc (Zn), and precious metals, silver (Ag) and gold (Au), from sulfide minerals [1,2]. An alternative method to smelting was initially sought to combat the emission of acidic sulfur dioxide ( S O 2 ) gas to the atmosphere and the loss of associated precious metals to slag [3,4]. Now, for copper, approximately 20% of total primary production is achieved via hydrometallurgical processing [5], and hydrometallurgy potentially offers advantages over smelting with respect to lower capital and operational costs, in the recovery of precious metals, recycling of process streams, and the production of inert, environmentally stable, waste residues. The advantage also of constructing a small-scale hydrometallurgy plant at the ore processing site is of increasing importance where transport costs are significant and for newly discovered deposits in remote locations [4]. Currently, commercial hydromet processes for copper largely feature sulfate-based heap or pressure leaching combined with solvent extraction and recovery via electrowinning, and are generally suitable for processing copper oxides or secondary copper sulfides, such as covellite (CuS) or chalcocite ( C U 2 S ) . Chalcopyrite ( C u F e S 2 ) , the most abundant copper mineral, is however difficult to leach and competing sulfate based leach systems are currently investigating fine grinding and addition of surfactants to overcome passivation effects and product layer diffusion obstruction [6,7]. Chloride leaching, however, has been extensively proven to extract >95% of copper from C u F e S 2 at temperatures around the boiling point and at atmospheric pressure [3,4,8]. Chloride leaching offers significant advantages for hydrometallurgical processing, in supporting high metal solubility, enhanced redox behaviour, and increased rates of leaching. In this system, both the cuprous ion, Cu +, which is sparingly soluble in sulfate systems, and cupric ion, Cu 2 + , as well as a host of other metal ions are stabilized through complexation with varying arrangements of chloride ions [3]. The result is that high 1 recoveries of both the primary metal of interest and associated precious metals can be achieved, by leaching at fairly modest temperatures, and thus has potential for treating complex ores and concentrates. Other attributes of a chloride system include the minimization of sulfate formation and the formation of predominantly elemental sulfur that remains in the leach residue, and that pyrite is generally left undisturbed in the leach solids [4]. Concentrated chloride solutions also induce high ionic activities that can play a beneficial role in the precipitation and separation of metal chloride salts [1]. The chemistry behind the chloride system is complex but it is this complexity that potentially allows for selective separation and recovery of valuable metals from process liquors or waste material. Yet despite the effectiveness of chloride leaching, there has been a definite lack of progression toward commercial activity. The reasons for this are largely due to problems with product purity and morphology, in addition to changes in the market value of copper and competition with existing smelting techniques [4]. In this thesis work chloride leach experiments are conducted with respect to two process flowsheets, shown in Figures 1 and 2. Both process models, feature chloride leaching of chalcopyrite combined with solvent extraction, an effective method for the selective removal of copper, and copper recovery by conventional sulfate electrowinning, to address the issue of pure product recovery. They differ however with respect to iron deportment. The hematite model, in which no air is introduced to the leach, allows separate sulfur and iron residues to be produced, thus making waste treatment and disposal potentially easier. The goethite model, however, offers simplicity in overall flowsheet design, and air is added to the leach to precipitate iron as goethite (FeOOH) for disposal with the leach residue. This work builds upon previous experiments undertaken at UBC [9] in which a number of experimental difficulties were encountered performing a four stage countercurrent leach, particularly with respect to S-L separation. For simplification this work has 2 reduced the number of leach stages to two and examined the use of a novel apparatus to aid smoother S-L transfer. Figure 1 The Hematite Process Figure 2 The Goethite Process CuFeS, 2 Stage Leach Residue (Fe 2 0 3 -hematite") Residue (Sulfur) O, Cu SX A/clave Fe removal - O , E W (from sulfate) T Cu° Air to LS2 CuFeS 2 • 2 Stage Leach Cu SX Residue (S & FeOOH-goethite) E W (from sulfate) T Cu° The goals of this investigation were: 1. to complete mass balances for both models to determine feed liquor requirements for leaching; 2. to perform continuous 2 stage countercurrent leach experiments for both models varying residence time, temperature and particle size to determine optimum leach conditions; 3. design and implement a bench scale, glass "mini-thickener" to improve countercurrent transfer of the solids and liquids 4. to assess speciation in the pregnant leach solution with respect to downstream processing. 3 2 L I T E R A T U R E R E V I E W 2.1 Aqueous copper and chloride chemistry 2.1.1 Chloride complexes One of the most significant features of the chloride system for copper hydrometallurgy is that the cuprous, C u + , ion becomes stable through complexation with chloride ions. In a pure water or sulfate solution C u + ions are not stable and the cupric, C u 2 + , ion is the dominant species [10]. In water at 25°C cuprous chloride solubility is of the order 2 x 1 0 " 7( however this solubility increases significantly with increasing total chloride concentration [11]. Figure 3[12] shows the relative thermodynamic stabilities of the copper ions in the C u - H 2 0 system at 25°C in a Pourbaix (Eh-pH) diagrams, and Figures 4[12] and 5[12] show the C u - C l - H 2 0 system at 1 M CI and 5 M CI, respectively. The ability to support both C u + and C u 2 + ions has major process implications for chloride systems with respect to the solubility limit of copper, the redox potential of the process liquor, and product recovery, particularly the decreased energy requirement to electrowin C u + in comparison with C u 2 + . The key issues for chloride complexation are complex stability, the comparative strengths of metal chloride complexes and the expected variations for metals in different chloride concentrations. For many metal chloride complexes, the maximum number of chlorides is usually four [13]. In the case of the cupric ion, substitution takes place with the four strongly bound equatorial water molecules [2], thus forming a complex with a square planar geometry [2,14]. 4 Figure 3 Cu-H20 E h pH Diagram at 25°C (1 atm, 1M) [12] Eh (Volts) 1 1 1 1 1 1 1 1 1 1 1 1 1 " _Cu20 Cu I I I I I I 1 I I I I I I 0 2 4 6 8 10 12 14 pH Figure 4 Cu-Cl-H20 E h pH Diagram at 25°C (1 atm, 1M) [12] Eh (Volts) 1 1 1 CuCl(+a) i i i i i 1 1 1 1 CuO CuCI ^ " " " l _ _ _ _ _ L____Cu20 - — _ _ _ _ _ _ -Cu i 0 2 4 6 8 10 12 14 PH 5 Figure 5 Cu-H 2 0 E h pH Diagram at 25°C (1 atm, Cu 1M, Cl 5M) [12] Eh (Volts) 1 1 1 1 1 1 1 1 1 1 1 1 1 CuO CuCI3(-2a) I ^ T u M ? Cu 1 I I I - I I 0 2 4 6 8 10 12 14 p H In general, metal ion complexes (MLn) are formed by the successive addition of ligands (L) , and the complex formation reactions can thus be written (a = activity) [10]. cl M z + + L" 1 <t> M L 2 " 1 K , = — ( 2 . 1 ) V + a L ' Thus for cupric chloride complexes the following exist: C u 2 + + Cl" « • CuCl + K, = A C U C 1 + (2.2) acu2+acr cucr + c r <=> Cuci2 K 2 = A C U C ' 2 (2.3) cl 1 ci Cuci+ cr CuCl 2 + Cl" « • CuClj K3 = A C U C L ; (2.4) a C u C l 2 a C l -C u C l j + C l ' o CuClJ- K 4 = A q v C i 1 ' (2.5) a a Cuci; cr 6 Kn=i_4 represent the stepwise stability constants for each separate ligand addition, but cumulative stability constants, p\, are also often quoted and are defined by: M z + + riL"1 « • M L z n " n yfti = M L ° " (2.6) a L -Many studies have been undertaken and data tabulated for the stability constants of metal-chloro complexes [2,10,11, 15-18]. This information allows for the determination of complex speciation for various concentrations of chloride and the proposal of relative strengths of chloro-complexes [10,11]. From these studies, the following order of strength has been suggested for metal-chloro complexes, from the strongest (Cl" acceptors): A g C l > C u C l > P b C l 2 > Z n C l 2 > C u C l 2 > F e C l 3 > F e C l 2 > N i C l 2 > HCI , N a C l , KC1 (Cl" donors) [11]. Table 1 [3] provides an overview of metal-chloro complexes and indicates which species are dominant with respect to changes in chloride concentration. Table 1 Metal-chloro complexes with increasing chloride concentration [3] Low Cl" concentration High Cl" concentration Cu (II) C u 2 + CuCf CuCl 2 CuCl 3" CuCl 4 2" Cu(I) CuCl 2" CuCl 3 2 " CuCl 4 3 " Fe (III) Fe 3 + FeCl 2 + FeCl 2 + Fe (II) Fe 2 + FeCf Zn Zn 2 + ZnCf ZnCl 2 ZnCl 3" ZnCl 4 2" Pb PbCf PbCl 2 PbCl3" PbCl 4 2" Ni N i 2 + N i C f Co Co 2 + c o c r Mn M n 2 + M n C f Cd C d 3 + CdCl + CdCl 2 CdCl 3" CdCl 4 2 " Sb SbCl 2 + SbCl 2 + SbClj SbCl4" SbCl 5 2" SbCl 6 3" Bi B i C l 2 + B i C l 2 + B i C l 3 BiCl 4 " B iCl 5 2 " B iCl 6 3 " As AsCl 3 Ag AgCl 2" AgCl 3 2 " Hg HgCf HgCl 2 HgCl 3" HgCl 4 2" 7 The formation of chloro-complexes is largely dependent on the chloride concentration, as can be seen by the above equilibrium equations, and tends toward higher substitution by chloride at higher concentrations [2,3,11]. For cupric chloride, C u 2 + , C u C l + and C u C k , and possibly CuCh," and Q1CI4 ", are observed with increasing chloride concentration, and similarly for the cuprous ion, C u C V , C u C ^ 2 " and CuCU 3 " , are observed. Also , polynuclear cuprous chloride complexes, i.e. CU2CI42", Cu 3Cl6 3", CU2CI3", have been suggested [3,16] as probable species and a spectrophotometric study of a highly concentrated C u C l - N a C l - H C l - H 2 0 system has most likely identified CU2CI42", rather than CuCl 4 3 "[3] . It is important to note that complexes may have a variety of charges (positive, negative or neutral) and that the chlorides may be replaced with other ligands, i.e. substituted, or the complexed species themselves may be complexed by other ligands, which can be useful in solvent extraction techniques [2]. The relative strengths of each of the metal ion complexes become important in chloride solution leach liquors. 2.1.2 Solubility It is important that the primary metal of interest for a certain process has an appreciable solubility to allow fast leaching and efficient movement through the process, particularly for economical reasons. Most of the metal ions important to copper chloride systems have greater solubilities than in sulfate solutions, e.g. C u 2 + (five times), F e 2 + , N i 2 + and Z n 2 + [1,20]. Particularly for the generally water insoluble chloride salts, i.e. PbCl2, C u C l and A g C l , much greater solubilities can be achieved in highly concentrated chloride solutions. For example the solubility of C u C l in cold water is 0.6 mg/L but in concentrated brine solutions >100g/L can be achieved (see Figure 6 [1]). The solubility of PbCL. in 25°C water is 1 g/L but a solubility increasing ten times can be achieved at 6 M total chloride (see Figure 7 [1]). 8 Figure 6 CuCl Solubility [1] 1501-o 100 5 50 L E G E N D -5°C 27°C 95°C FeCI, A 0 • +0-25 HCI 9 a +0-50 HCI • m +0-25 MgCI, C n + 0-50 MgCI, O a 2 3 4 5 6 Total C l " (Equivalents per l i tre) Figure 7 PbCl2 Solubility [1] 9 Though usually present in smaller amounts in the concentrate, A g recoveries have been high in most chloride extraction processes as it possesses a high solubility in strong brines at the moderately high temperatures used, e.g. above 85°C.[1] A s shown by the above two figures the effect of temperature is quite significant, with solubility increasing with higher temperatures. This effect has been used in the recovery of these metals in piloted chloride processes [3,4], where by simply reducing the temperature the chloride crystals, i.e. C u C l and P b C h , precipitate out of solution, with purity relative to the contents of the process solution. It has been shown that solubility limits vary not only with the total chloride concentration but also with the type of chloride salt added to the solution. The reason for this is that different metal cations form chloride complexes of differing strengths thereby changing the availability of chloride ions for complexation with typically water insoluble species. For example, Fe forms weak complexes in chloride solutions therefore leaving the chloride ions available to complex with C u + , and thus increases the solubility of the cuprous ion. Conversely, Z n 2 + ions form strong complexes with chloride ions and thus decrease the solubility of the cuprous ion [11]. In conclusion, for chloride hydrometallurgy it is then desired that a high background concentration of chloride salts (e.g. K , Na , M g , Ca) be present in the leach/process liquor at a moderately high temperature (e.g. near boiling point) for maximum extraction of copper and other valued metals. Despite the possible negative effect on copper solubility by other metal cations present, as long as a significant amount of chloride ions are present, impurities in the leach liquor are generally not a concern. Impurities however w i l l cause concern during product recovery stages. 2.1.3. Ionic Activity In dilute solutions, e.g. <1M, the chemical potential, or activity (a), o f a species can be approximated by the Molar concentration ( M , mol/L), or molal concentration (m, mol /Kg H2O). Therefore by the following relationship the activity coefficient (y) is 1 [1,13]. 10 Activity (a) = Activi ty Coefficient (y) x Concentration (m,M) (2.7) However in very concentrated solutions, e.g. >2M chloride (i.e. concentrations typically used in chloride hydrometallurgy), there is a significant effect on the activity of the species present (shown in Table 2 [1]). In summary, a very concentrated solution can greatly increase the activity (or chemical potential) of species present. A s an example, a very small amount of acid can be used to keep the p H very low in a concentrated chloride solution. Table 2 Mean activity (a) of H+and Clin 2M HCI and several salts [1] Salt Concentration N a C l C a C l 2 M g C l 2 _ 2.02 2J02 2.02 2 N 4.03 4.89 5.23 4 N 7.28 12.05 13.4 6 N 12.4* 27.50 30.6 * Supersaturated in N a C l In addition, cations that have higher charge densities (e.g. Mg 2 + >Ca 2 + >Na + ) and are present as undissociated species, have a higher hydration number, i.e. have a greater number of interactions with water molecules, and thus decrease the activity of the water and increase the activity of ionic species present. Wi th respect to the oxidative leaching of chalcopyrite in a concentrated chloride solution, the influence of chloride ion activity is quite important. For cupric leaching the final reaction may be stated as [ 1 ]. CuFeS 2 + 3 C u 2 + + 8C1- -» 4CuCl 2 " + F e 2 + +2S° (2.8) A n d at equilibrium, the equilibrium constant K is defined: 11 log K = 4 log a(Cuci2-)+ log ^ 2 + ) - 8 log a<ci-) - 3 log ^ 2 + ) (2.10) The aim of the leach is generally to achieve high cuprous and low cupric ion, i.e. to go to completion. This is thermodynamically favoured only at high chloride ion activities [1,10]. 2.1.4 Redox Chemistry The standard potentials for many species are altered in chloride solutions as a result of their chloro-complex formation (see Figure 8 [3]) [1,3]. The formation of strong complexes, e.g. C u + , P b 2 + and Z n 2 + , decreases the potential at which these metal species are reduced to their metallic state. This can be observed by using the Nernst equation (2.11) to calculate the reversible potential for the species MCl n z " n /M° in chloride solution, with respect to the standard potential for M z + / M ° at 25°C. F = F ° ° - 0 5 9 i l > c r m n Z a M C l -A t higher B, i.e. for stronger complexes, the deposition potential decreases. A s an example the potential for cuprous deposition decreases from 0.521 in water to 0.084 in concentrated chloride solution [3,13,21]. The advantage of maintaining copper in the cuprous state is that it allows deposition of copper via a one-electron transfer, i f direct electrowinning is used, and therefore offers potential energy savings when compared to cupric deposition, requiring 2 electrons per copper atom. Also , the theoretical cell potential is decreased relative to electrowinning C u 2 + in sulfate via reducing C u + at the cathode and oxidizing C u + to C u 2 + at the anode. In addition, the Cu + /Cu° deposition is faster, having only to transfer one electron, with an 12 exchange current density, j0 (rate of electron transfer), in chloride solution ( 5 M NaCl) determined to be 2400A/m 2 at 25°C, when compared to that of the Cu(U)/Cu° deposition in sulfate which can be between 20-200A/m [3]. Figure 8 Altered redox potentials in chloride solution [3] Reversible potentials in a 4M NaCl - Eh (mV) Standard potentials 0.5M HCI solution at 30°C 1600 - j 1400 -- 1500 Au(lll)/Au - 1360 eve r Cl 2 /Cr - 1242 1200 -Au(lll)/Aiu 1012 1000 -800 - Fe(lll)/Fe(ll) - 771- 779 Ag(l)/Ag Fe(lll)/Fe(ll)- 681 Cu(ll)/Cu(l) - 584 600 -- 521 Cu(l)/Cu Ag(l)/Ag - 335 400 • Cu(ll)/Cu- 337_ 320 Bi(lll)/Bi Bi(lll)/Bi- 9 1 - 84 Cu(l)/Cu 200 -o - 153 Cu(ll)/Cu(l) Ni(ll)/Ni - -58 u -200 - _ . 2 5 0 Ni(ll)/Ni Fe(ll)/Fe - -458 -400 i -600 --800 -_ -441 Fe(ll)/Fe Despite the ability to deposit copper at a fast rate at the cathode, and although this is good for process efficiency, the downside is that the copper is not smooth and fine metallic crystals of copper are formed instead. It should also be noted that the potential ohmic drop for chloride solutions is larger than in sulfate, i.e. the loss of energy as current travels through the electrolyte is greater. 13 The growth of finer copper crystals (e.g. dendrites, nodules or metallic powder) may lead to greater contamination due to the exposure of a greater surface area to the process liquor, and generally these morphologies also make the product harder to handle. Investigations have consequently been made into the use of additives to aid in better, smoother, product formation [3,16]. However, greater product purity problems arise from the presence of more noble elements in the cell feed solution, e.g. selenium (Se), tellurium (Te) and most notably silver (Ag), and these w i l l even in small concentrations deposit at potentials required for C u + deposition, as this potential w i l l be below the reduction potential for these elements [4]. O f particular interest for gold bearing concentrates, is the significant decrease in the standard potential for the A u 3 + / A u ° couple in concentrated chloride solutions, from 1.5 V to 1.012V in 4 M N a C l - 0 . 5 M H C I (Figure 8) [3]. However of overall importance is the effect of the concentration of chloride on the oxidation potential of the couples Fe and C u , both of which are commonly used in the oxidative leaching of copper sulfide minerals. The standard reduction potentials (at 25°C) for F e 3 + / 2 + and C u 2 + / + are 0.77 and 0.153, respectively, and therefore the ferric ion has a far greater oxidizing potential in pure water solutions [3,22]. In 4 M N a C l , 0 .5M H C I solutions the oxidizing potential of the iron ion couple falls to 0.681 and the copper ion couple increases to 0.584, both of these potentials are high enough for oxidizing most metal sulfides [3]. These changes are again due to the complexation behaviour of the ions. For copper in chloride solution the electron transfer from C u 2 + to form the reduced C u + species, i.e. C u 2 + + e <=> C u + , can be more correctly written as: C u 2 + + 2C1" + e 1 _ <=> CuCLf. Cuprous is more strongly complexed than the cupric ion, which forms weaker complexes and remains largely uncomplexed to high chloride concentrations. Therefore the ratio of free C u + : C u 2 + decreases with increasing concentrations of chloride and according to the Nemst equation the potential of the C u 2 + / + couple w i l l thus increase [13,21]: 14 RT a E ^ c ^ ^ - T l n - ^ (2.12) r d C u 2 + Conversely, for the Fe /Fe couple the oxidizing potential decreases with increased chloride concentration as the ferric ion forms stronger complexes with the chloride ions than does the ferrous ion. Therefore the ratio of free F e 3 + : F e 2 + decreases with increasing chloride concentration and thus the solution potential decreases. Similarly the relationship can be observed [13,21]: E=K*^-^-^ (2-13) Figure 9 [23] shows the effect of chloride activity on the potential of species common to copper chloride hydrometallurgy solutions. Figure 9 Eh vs Log a a [23] - 2 - 1 0 1 2 log cr 15 The importance of a concentrated chloride solution for the oxidative leaching of chalcopyrite by cupric is that a high oxidation potential can be maintained as the leach progresses. A high ratio of cuprous, reaction product, to cupric can be present, and the reaction still be driven to near completion. This is because cuprous is associated only as chloro-complexed species, effectively reducing the free cuprous ion concentration. 16 2.2 Chalcopyrite leaching in chloride 2.2.1 Thermodynamics and kinetics To determine the suitable aqueous conditions in which leaching of a particular mineral can be carried out, an Eh-pH (or Pourbaix) diagram assists by indicating the predominant species, of metal and metal compounds, present at equilibrium with respect to reduction potential and hydrogen ion concentration. B y following the progression from mineral to soluble metal species on the diagram leaching pathways can be determined, i.e. what adjustments in p H and/or solution potential are required to liberate the metal ions contained in the mineral [22]. Eh-pH diagrams represent only thermodynamic equilibria however and so alone do not indicate which leaching pathway w i l l work best. The kinetic factors of each reaction must also be considered and are especially important when considering process design. This kinetic information comes from experimentation and observation, and leaching rates are generally influenced by reactant concentration, temperature, particle size (surface area) and product layer formation. In this study the copper sulfide mineral chalcopyrite (CuFeS2) is of particular importance as it is presently the most abundant source of copper [22]. B y observation of the Cu-Fe-S-H2O Eh-pH diagram at 25°C, which has been simplified to include only laboratory observed reactions (See Figure 10 [24]), it is also confirmed to be quite refractory, being quite resistant to acid attack and requiring strong acid oxidizing conditions [6]. The stability of chalcopyrite has been attributed to its face-centered tetragonal lattice structural configuration [25]. 17 Figure 10 C u - F e - S - H 2 0 E h - p H diagram at 25°C [24] 18 Figure 11 C u - F e - S - C l - H 2 0 Eh-pH diagram at 25°C [24] 19 A s this study is also focused on chloride solutions, Figure 11 [24] shows that when chloride (at 1M) is added to the Cu-Fe -S -H 2 0 Eh-pH diagram at 25°C the regions of sulfide stability are not significantly affected and the only difference appears to be in soluble copper and oxide regions. Thus during leaching the presence of chloride appears to prevent oxide formation, that could possibly slow the leach reaction [24]. A s a strong acidic oxidizing solution is required, studies and processes for chalcopyrite leaching [1,3,4,10,21, 26-33] have so far employed ferric and/or cupric chloride and have successfully demonstrated the effective leaching of copper at atmospheric pressure in solutions near boiling point. Consequently, the following two leaching reactions are generally accepted: CuFeS 2 + 4 F e C l 2 + -» C u 2 + + 5Fe 2 + + 2S° + 8C1" (2.14) CuFeS 2 + 3 C u 2 + + 8C1" -» 4CuCl 2 " + F e 2 + + 2S° (2.15) Not predicted by the Eh-pH diagram is the recovery of predominantly elemental sulfur (S°) instead of the oxidized sulfate species. In fact less than 5% of the S in chalcopyrite generally oxidizes to sulfate in atmospheric chloride leaching, and it is when pressure leaching is employed that greater sulfate formation may become a concern [34]. The formation of a S° product layer around the mineral particles has not been observed to affect the rate of leaching [34], and more discussion regarding the fate of sulfur in chloride leaching is presented in a section 2.2.2. To further complicate the reaction kinetics and mechanisms involved, and despite the reporting of separate studies for ferric and cupric leaching, the two oxidant species are closely linked as ferric leaching w i l l produce cupric ions, cupric leaching w i l l produce ferrous ions, and ferric ions w i l l oxidize cuprous ions to form ferrous and thus regenerate cupric for further leaching, as in the following reaction: F e C l 2 + + C u C l 2 " -» F e 2 + + C u 2 + + 4C1" (2.16) Before discussing the findings of kinetic leaching studies it is important to note that, by simple observation of the above reactions, there is flexibility in which form the copper 20 ion may be attained during or at the end of the leach, and this may ultimately be important for downstream processing. Also due to the role chloride plays in stabilization of the cuprous ion, copper solubility and the C u 2 + / C u + redox potential, chloride concentration is to be a significant factor in the leaching rate and extent of leaching. It was discovered early on that the ferric oxidation of chalcopyrite in chloride solution was much faster than in sulfate solution [21,27]. However it was not until the development and testing of hydrometallurgical chloride processes for copper in the early 1970s [3,4] that there was significant interest in determining the reaction mechanism and kinetics behind leaching in chloride media. It was then found that the presence of cupric substantially increased the rate of leaching [21,32]. Table 3 below summarizes the main findings of chalcopyrite leaching studies in chloride media and following the table is a brief discussion on important leaching aspects. The two requirements that enable optimum leaching with cupric ions (Equation 2.15) include the presence of a significant concentration of chloride and also that the potential is low enough to support the cuprous ion [10,21,32]. However, as stated previously, in a chloride solution the C u 2 + / + redox potential is raised due to chloro-complexation, and therefore the potential required for cupric leaching does not have to be too low, provided the chloride concentration is high. It has also been reported that to leach to completion, a high chloride concentration, high temperature and short leaching time must be employed to lessen the reduction of elemental sulfur by the cuprous reaction product (Equation 2.17) [30]. 2 C u C l 2 " + S ° -» C u S ( s ) + Cu 2 + +4Cr (2.17) 21 Table 3 Summary of chloride leaching kinetics Linear kinetics observed [10,21]: Extraction rate was shown to increase with temperature; this is an indication of surface rate control (i.e. rate controlled by electron transfer across mineral surface) as reaction products do not affect leaching rate; activation energies (>40kJ/mol) determined from temperature experiments also indicate surface control of dissolution rate. Temperatures near boiling point are therefore best to achieve maximum extraction, higher temperatures would cause sulfate formation (See Section 3b.). Smaller particles increase rate [10,21,28]: Following from the above, as the surface is rate controlling, an increase in surface area would therefore increase the rate of particle dissolution, and this has been demonstrated. Rate dependent on [10,21,28]: Oxidant concentration (i.e. cupric or ferric), cuprous and chloride concentration; also the cupric leaching rate has shown to be greater than ferric. Rate independent of [21,28]: Hydrochloric acid concentration, ferrous concentration and completely dissociated chloride salts; p H should however remain low enough to avoid the precipitation of copper-oxy chloride or iron compounds; rate is also independent of stirring rate, which needs only be sufficient to suspend particles. The presence of air, or stronger oxidants, in the leach can assist also in oxidizing the cuprous ion product to subsequently prevent cuprous reduction of sulfur and/or regenerate cupric for further leaching. Oxidation of cuprous by air may occur either directly (Equation 2.18) or via the oxidation of ferrous to ferric (Equation 2.19) which may then catalyze the oxidation of cuprous to cupric (Equation 2.20). 2 C u + + 2 H + + 0.5O 2 -> 2 C u 2 + + H 2 0 (2.18) 2Fe 2 + + 0.5O 2 + 2 H + -> 2Fe 3 + + H 2 0 (2.19) C u + + F e 3 + C u 2 + + F e 2 + (2.20) 22 The conversion of ferrous to ferric may also allow for the removal of iron as goethite at ambient pressure during the leach process via the following reaction (see section 2.2.3 on iron deportment): F e 3 + + 2 H 2 0 -» 3 H + + F e O O H ( s ) (2.21) It is not always desirable for the iron to be removed with the leach residue and so to prevent the above from occurring oxidant addition needs to be controlled. Despite the possible array of proposed intermediary Cu/S compounds, from the reaction of chalcopyrite with either ferric and/or cupric ions, generally in concentrated chloride solutions there is observed to be a simultaneous rapid extraction of both copper and iron. This indicates that potentially hindering reactant products are not formed for either species. In summary, there are a number of factors that contribute to the rate of CuFeS 2 leaching and an atmospheric leach at boiling point, low p H , with relatively fine particles and a high concentration of both oxidant and chloride are important for efficient and effective leaching. 2.2.2 Sulfur Both fundamental studies and developed hydrometallurgical processes involving chalcopyrite leaching with F e C i 3 or C u C l 2 have generally reported >95% of the sulfur reaction product forming as elemental sulfur, and hence <5% oxidation of sulfur to sulfate [4,34,35]. Even when long retention times (up to 90hr) were employed, or large amounts of oxidant or acid were used, at temperatures below 100°C there was still found to be predominantly elemental sulfur as the sulfur leach reaction product [34]. The advantages in the recovery of sulfur in its elemental form include the following [4,34,35] fugitive emissions of S 0 2 are avoided, unlike in smelting; easy storage and handling; 23 limited environmental hazards (in waste residues sulfate hydrolysis causes acidic conditions in the environment); oxidant is consumed only by the desired mineral reaction, thereby maximizing the release of the metal of interest (i.e. no oxidant is wasted on S oxidation); the precipitation of sulfate containing jarosite is avoided (Jarosite is typically a process waste residue for the removal of iron and tends to be less stable in the environment- see section on Fe deportment). However, i f so desired jarosite may be formed with the aim of removing sulfate from the process liquor; and, i f sulfate forms the leaching rate of the chloride system w i l l be much slower. Generally, sulfate forms when temperatures are employed above the melting point of sulfur (119°C) [35]. This explains the increased sulfate levels (up to 25% of sulfur product) of chloride hydrometallurgy processes which involve pressure leaching, e.g. in the C L E A R process. However, although for relatively few processes, 5-25% sulfur oxidation has still been reported for systems that remain below 100°C and it has been suggested, that although not precisely clear, it is likely that the mineralogical composition of the concentrate tested was a factor [4]. For example some sulfate may originate from the superficial oxidation of the sulfides prior to leaching (e.g. PbS04 on galena, PbS) or by the sulfide leaching reaction itself. In addition, it has been shown that despite the presence of air in the leach there is no significant effect on the relative amounts of elemental sulfur and sulfate produced and the relative amount of elemental sulfur formed is independent of chalcopyrite particle size [34]. The most likely mechanism of elemental sulfur formation appears to be via the dissolved intermediary reaction product H 2 S . This mechanism is reasoned by the presence of wel l -formed euhedral sulfur crystal growth that indicates the elemental sulfur was deposited from aqueous solution, as opposed to conversion of the sulfide remaining on the mineral surface. A variety of morphological arrangements of these sulfur crystals have been observed (dependent on mineral particle size and preparation) ranging from large globules to continuous films. However, due to observed linear leaching kinetics 24 sufficient porosity must exist in this product layer to allow passage of reactants and thus does not cause passivation [34]. One advantage of the sulfur coating of particularly fine CuFeS2 particles is to cause agglomeration thus improving thickening and filtration properties of leach residues [4]. The addition of calcium to process liquor, as either a chloride salt (CaCL:) or carbonate (Limestone, CaC03, typically employed for p H adjustment) is a very good method of removing sulfate, as gypsum (CaS04), from solution. Jarosite may also be precipitated at moderate leaching temperatures and in acid concentrations but this reaction w i l l still leave a background concentration of sulfate (e.g. lOg/L) [4]. 2.2.3. Iron Deportment Iron is the fourth most abundant element in the earth's crust and as a consequence is usually present in most metal ores, including those of copper. Iron may be present as an essential constituent of the ore mineral (i.e. the material containing the metal of value), e.g. chalcopyrite CuFeS2, bornite Cu5FeS4 and cubanite CuFe2S3, and therefore in extracting the metal of interest the associated iron w i l l also accumulate in the leach liquor, and w i l l have to be removed prior to recycling of the stream for leaching. Iron may also be present in gangue minerals (i.e. materials that do not bear metals of value for the process) either as an essential constituent or partially substituted for an essential constituent. The most common iron-bearing gangue mineral in sulfide ores, and in particular for the chalcopyrite ores this study is interested in, is pyrite FeS2. For oxide ores the most common include hematite Fe203 and magnetite Fe 3 04. In addition, owing to iron's similarities in chemical properties with many elements (in particular the first series transition metals) it may substitute for an essential element into the crystal lattice of either ore or gangue mineral [36]. Whilst it is generally unavoidable to accumulate iron in the leach liquor, i f iron is present in the ore mineral, it is possible to avoid dissolution of gangue materials by selective 25 leaching conditions. For example a chloride leach for chalcopyrite operating at atmospheric pressure enables this strategy as pyrite is generally left unattacked. Pyrite in fact reacts much more slowly than chalcopyrite even in oxygen pressure leaching (especially <170°C [36]) and this may be due to the formation of elemental sulfur around the pyrite and/or galvanic protection by less noble minerals such as galena, covellite, and sphalerite [22]. Iron still however makes up an appreciable content of the chalcopyrite leach liquor owing to its 1:1 Cu:Fe ratio, and the most common method for removing iron from hydrometallurgical processes is via the precipitation of iron compounds. The precipitation can occur during the leach stage, and therefore iron is removed with the leach residues, or separated at another stage of the process, which may be preferred so as not hinder the removal of sulfur or precious metals from the leach residue or for the production of a separate saleable iron product. The three predominant iron compounds that are precipitated in hydrometallurgical processes are goethite F e O O H , hematite Fe 2 03, and jarosite (H 30,K,Na,NH4).Fe3(S04)2(OH) 6 compounds, i f sulfate is present [36]. The precipitation of these iron compounds occurs when the iron is present as ferric ion and there is low acid. When oxygen is added to convert ferrous to ferric, as seen by the following equation, then this consumes acid, and can thus raise the p H to initiate precipitation. 2 F e 2 + + 0.5C-2 + 2 H + -» 2 F e 3 + + H 2 0 (2.22) If the p H is raised too quickly, with poor agitation at low temperatures [22] ferric hydroxide w i l l form and due to the polymeric nature of this gel colloid precipitate, which contains cross-linked hydroxyl bridges, it is poorly filtered and w i l l not be successfully separated from the process. On the other hand a steady p H increase, high temperatures and good agitation w i l l result in the well-structured, easily filtered iron compounds, 26 goethite, hematite and jarosite. In a chloride leach stream the iron is often produced as ferrous and therefore by careful mixing with oxygen, to convert ferrous to ferric, the rate of p H increase can be controlled, and therefore a filterable iron compound is easily achieved. Table 4 [36,37] summarizes process requirements for precipitation of these compounds, with comments on their stability as waste residues in the environment, and their role in hydrometallurgical processing. It is important to comment on one study [37] that showed by increasing the chloride content of solutions with chloride salts that completely dissociate, e.g. N a C l or CaCb;, the proton activity increased, thereby decreasing the quantity of hematite produced, without however effecting the composition of the product. When salts which formed complexes with chloride, e.g. ferrous or zinc ion, were added there was no effect of the amount of hematite formed, as the free chloride did not change and therefore the proton activity did not change. 27 Table 4 Summary of i ron compound precipitation [36,37] Iron Compound Process Requirements Comments Hematite F e 2 03 (Fe ~ 70%) Generally formed at T>130°C in an autoclave (T>100°C i f hematite seed is present). When formed in chloride solutions C l found to be <1% of product. Though more costly to produce, it is very stable and good for ponding and possible resale. Stability demonstrated in natural environment, by its occurrence as a weathering product. High Fe% requires less solids to be handled. A study has shown also that by seeding precipitation a more coarse grain is formed which is easier to filter. Goethite F e O O H (Fe ~ 63%) Generally formed at T<120°C. When formed in chloride solutions product may contain - 8 % C l . Precipitation can occur at atmospheric pressure. Goethite has also been shown to be environmentally stable, after hematite, particularly in wet conditions. In fact the weathering of hematite in a humid environment usually produces goethite, 2FeOOH <=> F e 2 0 3 + H 2 0 . Jarosite M * F e 3 ( S 0 4 ) 2 ( O H ) 6 M * = H , Na , K , N H 4 + Fe ~ 35%) Precipitates over a range of temperatures and is dependent on the content of sulfate. Can be formed at a lower p H . The iron content is low. Thus more product is required to remove a comparative amount of iron. There are problems with long term stability as acid w i l l leach out: N a F e 3 ( S 0 4 ) 2 ( O H ) 6 + 3 H 2 0 -> '/JNfoSCU + 3Fe(OH) 3 + 3 / 2 H 2 S 0 4 28 2.3 Overview of chloride leach processes for copper Over the last thirty years a number of hydrometallurgical processes that employ chloride leaching for copper have been tested at the laboratory, pilot plant and/or demonstration plant [2,4]. These processes have been quite successful in exhibiting efficient extraction of copper, via leaching at atmospheric pressure and boiling point, but only one process, the Duval Corporation's C L E A R process, has been operated commercially. This process operated for a period of six years until it was closed due to various technical and economic reasons, noting that their product still required further costly refining to achieve wirebar- or cathode-grade material [4,38]. Product impurity (especially due to silver contamination) has in fact been a major hindrance towards commercial development of these processes along with technological constraints faced in the handling of corrosive process liquors [38]. Figure 12 provides an overview of the various methods of recovering copper from chloride leach liquor that have been investigated, and includes the chloride leach process proposed in this thesis work. Tables 5 and 6, on the following pages, summarize the leach and process characteristics of these processes, which were mostly designed to leach chalcopyrite (CuFeS2) concentrates. 29 Figure 12 Chloride leach processes for copper Chloride Leaching Recovery from crystals Electrowinning from chloride Solvent Extraction • C Y M E T [3,4,39] - U.S. Bureau Mines [i,4,13,21,24] UBC-Cominco [1,4,24] - Duval C L E A R [1-4,21,40] • Outokumpu [39] - Elkem [3,38] - Intec Copper [41,42] G C M [4] ' Electrowinning from chloride Electrowinning from sulfate Cuprex C M E P [2-4,38,43,44] - Minemet Recherche [45,46] - Phelps-Dodge [47] - Henkel Corporation [4,48,49] - *Falconbridge Copper Chloride Process [9] 30 » u < w u (Z3 w u o S O <D 60 03 .3 o i os CD S O CCS <D S B e S •+-• c o 00 o o OS a s os os as os ^3 so J3 so os OS 00 OS SO A oo ON u o H OS oo os so o o o o o o OS o o ON I o oo OS >/-) OS I </-> oo to O c3 U U u fe U C 3 3 u u i—) u ca CN 3 u a cd fe u C3 U 3 u <D fe c3 3 o fe u C3 3 u OH u C3 fe u C3 u + |3 Q OH fe u 3 U o i s u 3 u a CD 03 a. U <D fe to « o w U cs U 3 u CD fe H ha S 1 s £ vi >. CU -s o ffl 3 W u 3 u u 3^ u a) so O Q CD fe I CD "3 CD fe H 3 II u Q o Q <U £ • C/> <U O O i -CL| IT; CU H Z W o u H Z W o w Q Z o pi a PH > PH O o u U u 05 PH w H (/i oo a u o Pi PH u w u o PS" PH ii -a o g u s 3 o oo o A a b PH 3 U u u Pi UH U ca ^ oo| oo c oo N O J U * 3 oo < Pi U A 1-5 o d 3 u u 3 u Pi u A B cu cu ID PH O PH a Cu 6 o •a o ca 3 a d 3 u o 3 u U V £ O o c3 as 3 D •4-* c3 3 O Cu o 3 u 00 w A u o o oo a d u pa A Cu Cu <u PH 13 O PQ oo O Cu o 3 u a 00 < CN 00 .a 1-5 o a" d PH u pa v o 3 u u 3 w u 3 u oo u o pa Pi o u w X oo a N U H-> PH PH 6 3 u 3 o CN 00 c3 00 m .3 K 6 pa Pi 00 V pa v "co o 43 T3 3 oo .3 o c3 .3 O d u TH-IS o Cu o 3 u PH X 00 l-l u O O PH 3 u o I PH 3 u •B o o B PH 3 o B O O 3 PH < O 00 uJ §5 .a i-s d 9 U o pa u pa u X 00 pa o 00 X 00 pa o oo X 00 pa o 00 X 00 X 00 PH <: v w dp 00 00 t> •< pa V V o u pa <u PH o" U u 60 HO PH 3 N X oo V B Cu Cu O PH PH CQ X 00 Cu Cu 3 u J3 o ca — 'o < a u ca Cu u x 00 u u. u 00 O Q J -PH J3 o ca d o X 00 A a b PH pa o oo X 00 X 00 — R H >, u 03 £ S 3 & "> 41 •3 o CJ co (U UH PH CA CU CJ o ha OH hu o 3 T3 3 o p. a. o U o 3 C/3 —1 CQ 1 3 •c CN x m c o •a O CJ O oo £ "E c Table 5 shows process variations investigated for CuFeS2 leaching in chloride with, employing and/or ferric in a background of N a C l or CaCb., and these have all reported high extractions. Similar lixiviant characteristics and leach conditions were employed in this thesis work. Regarding the variations in copper recovery for each of the different processes, initial aims to crystallize pure crystals of copper salt, as cuprous chloride, found that silver was a major contaminant. Even when metallic copper was added to reduce cupric to cuprous, and cement noble impurities [1,4], this technique was not successful in completely removing the silver. Also , the temperature adjustments required and handling of the crystals may have potentially led to technological problems during process scale up. A s an example, evolved HC1 must be kept dry to prevent corrosion. Later attempts to electrowin copper from the cuprous state in chloride solution, attempting to reduce operating costs via electrowinning C u + , encountered the same product purity problems with respect to silver ( U S B M , Intec, Elkem, C L E A R ) [3,4,38,40]. These processes were also susceptible to product contamination by insoluble compounds as well as other noble impurities, such as antimony [13]. In addition, due to the high rate of winning in a chloride medium, the product was fine, either dendritic or powder-like, and thus created handling difficulties. Solvent extraction so far appears to be the best method of selectively removing copper from the leach solution, for preparation of a pure cell feed for electrowinning from either a chloride or sulfate solution. Organic extractants are generally highly selective for the cupric ion [49-51] and impurity transfer along with the copper is therefore minimal. However, it is important to wash the chloride from the loaded solvent to prevent transfer of chloride to the sulfate electrowinning cell, which may even in small quantities lead to corrosion [49-51]. Electrowinning copper in chloride solutions from the cupric state seems to negate the advantage of using chloride. Electrowinning cuprous, which is only stable in concentrated chloride solutions, would require only one electron and therefore half of the energy required to reduce cupric to metallic copper. Also partial and incomplete reduction of cupric to cuprous may occur. Electrowinning copper from sulfate solutions is a well-developed technology for producing smooth, high purity copper cathodes, and conventional cell designs are now well established. 33 However, commercial development of these chloride leach/sulfate electrowinning processes has likely been hindered by process complexity and uncertainty regarding the capital and operating costs required to compete with existing sulfate based leaching plants. The flowsheets of the three chloride leach processes in Table 6 that have proposed copper recovery via solvent extraction and sulfate electrowinning have flowsheets which differ with respect to this thesis work. The closest design is that of the Minemet Recherche process, in which the pregnant leach solution is split, one half going to solvent extraction and the other to atmospheric goethite precipitation, and then both streams combine to return to the leach. 2.4 Summary The two process models developed in this work combine chloride leaching with solvent extraction and electrowinning of copper from sulfate solution, and evolved from reviewing the various chloride processes previously attempted, see Figure 12. Although the experimental work in this thesis focuses on the leaching step, in hydrometallurgy it is important to consider downstream processing steps and recycling of the liquor to the leach to determine the suitability of a process. Mass balances were completed for both circuits, and leach experiments were performed with respect to the calculated feed liquor requirements. The leach conditions chosen for the countercurrent leach experiments were based on recommendations in the literature investigating the kinetics of chalcopyrite leaching. The literature indicates that an atmospheric leach, near boiling point, at a low p H , using relatively fine particles and a high concentration of chloride and oxidant are important for effective and efficient leaching of chalcopyrite. A s the reaction is linear with temperature and the rate determined to be surface reaction controlled, particle size, temperature and leach time were varied in the experiments to achieve higher extractions of copper. 34 3 EXPERIMENTAL 3.1 Mass balance spreadsheet preparation for the leach models For the two chloride leach process models considered in this thesis work, the "hematite" (no air in leach) model and the "goethite" (air in leach) model, simplified'mass balance spreadsheets were developed in Excel to determine recycle feed liquor requirements for the leach. These mass balances considered the leach with respect to downstream purification and recovery steps. In particular, the concentration of oxidants cupric and ferric were determined to allow 100% extraction of copper and iron from CuFeS2 and enable recycling of the process liquor to the leach following copper recovery. These recycle concentrations provided the basis for preparing feed liquors for continuous countercurrent leach experiments. Assumptions for performing the mass balance included identification of reactions taking place in each process unit and that 100% of the leach liquor moved countercurrently to the leach residues. 3.2 Feed concentrates The copper concentrate tested in all but the final experiment originated from Antamina in northern Peru, and 30kg was shipped via Falconbridge Ltd. For the final experiment chalcopyrite concentrate from the Rosario deposit at Collahuasi in Chile was tested, and 4kg was provided by Falconbridge. Both concentrates contained chalcopyrite (CuFeS2), with differing amounts of gangue material, and the Rosario concentrate also included chalcocite (CU2S). To obtain a representative sample of the Antamina concentrate, for the leach tests, the complete concentrate sample was mixed and divided using an automatic splitter in the U B C Department of Min ing Coal Lab. The splitter was then used to produce smaller packages, e.g. 2.5kg lots, suitable for testing. The Rosario concentrate was sent for grinding and then divided equally on a bench to select a representative sample for testing. Prior to use in the leach experiments both concentrates were air dried, over approximately one week, and rolled with a stainless steel rolling pin to break up clumps. For some of the experiments in which finely ground concentrate was employed, a portion of both concentrates was sent to Process Research Associates (PRA) in Vancouver to perform grinding to 20-35 25microns. The final and only experiment that employed Rosario concentrate had undergone fine grinding. Initial characterization of the concentrates involved analysis of 30 elements (including C u and Fe) by Inductively Coupled Plasma (ICP) spectrometry (via aqua-regia digestion), total sulfur analysis and Si02 analysis. These analyses were performed by International Plasma Laboratories (LPL), a contract laboratory in Vancouver. Gold analysis was also performed by ChemMet Consultants in Vancouver. The results of these analyses can be found in Appendices 1 and 2 and a summary of the C u , Fe and S% are shown in the following table for the three experimental concentrate feeds. The 3 feeds include the as-received coarse Antamina concentrate and the Antamina and Rosario concentrates that had undergone fine grinding. Table 7 Cu, Fe and S % in feed concentrates Feed Concentrate Cu % Fe % S% Antamina (as received) 28.8 28.4 32.1 Antamina (fine grind) 28.5 28.7 , 32.4 Rosario (fine grind) - 26.8 26.8 35.5 Particle size analysis included wet sieving for the as-received coarse material, as a significant proportion of particles were larger than the smallest sieve size of 38um. For analysis of the finely ground concentrates a Malvern Mastersizer laser analyzer was used to assess particle size distribution. The Mastersizer is an instrument that determines particle sizes via the detection of light refracted by the various sized mineral particles, and is well suited for particles <38um. The coarse material was also tested using the Mastersizer to provide a comparison. Both particle size analyses were performed in the U B C Department of Min ing Coal Lab and the results are shown in the following table. Reports for the Mastersizer analyses can be found in the Appendix 3. 36 Table 8 Feed concentrate particle size Feed Concentrate Method Results Antamina (as received) Wet Sieving Malvern Mastersizer p90 = -75pm p50= 42pm, p90=106pm Antamina (fine grind) Malvern Mastersizer p50= 16pm, p90=41pm Rosario (fine grind) Malvern Mastersizer p50= 16pm, p90=37pm p50 or p90 = size under which 50% or 90% of particles exist Characterization of both concentrates also included x-ray diffraction spectrometry ( X R D ) , and was performed in the U B C Geology Department. The X R D spectra have been included in Appendix 4. The Antamina spectrum reveals that the concentrate consists largely of chalcopyrite with a low-moderate amount of gangue material consisting of pyrite, silica, sphalerite, zinc oxide and molybdenite. The Rosario spectrum reveals that the concentrate consists of two copper minerals chalcopyrite and chalcocite with a significant amount of gangue material, largely of pyrite and lesser amounts of silica, molybdenite, and sphalerite. 3.3 Feed solutions Feed solutions were prepared in deionised water (DI water) using technical grade ferric chloride (FeCl 3 ) powder and 4-14 mesh calcium chloride (CaCl 2 ) pellets obtained in bulk quantities from Anachemia Science. Laboratory grade cupric chloride dihydrate (CuCl2.2H 20) was obtained from Fisher Scientific or Aldr ich and concentrated hydrochloric acid (HCI) from Fisher Scientific. To achieve the reduced ferrous solution for the goethite model feed liquor powdered iron was used to reduce ferric chloride. Feed liquor concentrations were based on the mass balance outputs of the model. For both models, an approximate background concentration of free acid (as HCI) was aimed at around 3g/L, either to approximate the expected free acid concentration in liquor recycled from hematite precipitation in an autoclave or to keep acid sufficiently low to promote goethite precipitation. Total chloride concentration was aimed at greater than 5 M with additional free chloride contributed by either 1 lOg/L or 220g/L C a C l 2 . 37 3.3.1 Feed liquor for the hematite model To prepare the hematite model feed liquor ferric chloride was added first to generate heat to then aid dissolution o f cupric and calcium salts. Finally, hydrochloric acid was added and, to determine the actual free acid concentration, an undiluted sample was sent to IPL for analysis via oxalate masking N a O H titration. Following the reporting of two anomalously high free acid concentrations by IPL, additional free acid titrations were performed at U B C and for all other experiments the results were comparable. Appendix 5 contains the Free A c i d method used for the U B C titrations and shows a table comparing IPL and U B C titration results. 3.3.2 Feed liquor for the goethite model The feed liquor for the goethite experiments contained iron as predominantly ferrous. To prepare the liquor, ferric chloride and calcium chloride were first dissolved in DI water and then powdered iron added to achieve the desired ferrous concentration, according to the following reaction (3.1). 2 F e C l 3 + Fe° -> 3 F e C l 2 . (3.1) A higher calcium chloride concentration, of 220g/L, was used, following the first experiment, to make up for the reduced amount of metal chloride salt added. Following completion of the reduction reaction, indicated by stabilization of the redox potential at 450mV A g /Agci (at room temperature), cupric chloride and hydrochloric acid were added. The final solution was then titrated with acidic certified cerium sulfate solution (0.1M) to determine total reduced iron, and the method used can be found in Appendix 5. The ferric concentration was thus calculated as the difference between the total iron and reduced iron. 3.3.3 Feed liquor analysis In addition to free acid analysis, two diluted samples of the feed liquor (d.f.xlOO), one with 2% HC1 and one with only DI water, were prepared and sent to IPL for ICP 30 element analysis and total chloride ion analysis respectively. Total chloride was analyzed by P L using an ion selective electrode (ISE) and reported as g/L HC1. Certificates of analysis for the IPL results can be found in Appendix 1. 38 3.4 Experimental Apparatus The layout of the two-stage countercurrent leach circuit is shown in Figure 13. For solid-liquid (S-L) separation one mini-thickener (Thl) , was incorporated between Leach Stage 1 (LSI) and Leach Stage 2 (LS2) and pressure filtration (P.F.) was used to separate final leach residues. The filtrate from pressure filtration (LS2 liquor) was then transferred countercurrently to L S I . The equipment and apparatus used is detailed in the following sections. Figure 13 Experimental Set-up Concentrate Feed Exit PLS LSI Fresh Leach Solution Leach liquor to LSI LS2 Solids Leach Residue - Final Filtration 3.4.1 Leach vessels The experiments were conducted using 2 L glass water-jacketed leach vessels, approximately 11cm in diameter, for containing 1L leach volumes for Leach Stages 1 and 2 (LSI and LS2). These tanks had indents, to act as baffles, and were sealed with a removable high-density polyethylene (HDPE) l id. The l id, important to prevent evaporation loss, had several stoppered openings for the temperature probe, p H and E measurement, sampling, stirrer shaft and connection to a condenser. Water condensers were used to maintain atmospheric pressure and prevent water loss due to evaporation. 39 A s the lixiviant, containing ferric and cupric in a total chloride concentration >5M, was extremely corrosive and corrodes stainless steel, equipment was largely made of either glass, titanium, or in some cases plastic was used (e.g. the lids and tubing). The titanium stirrers were made of grade 2, 8mm (5/16 inch) diameter rod, featuring a 6 blade radial design, and were attached via overhead motors. The stirrer was suspended approximately 2.5cm from the bottom of the leach vessel, which was 25cm deep, and each radial blade was 2cm long. The stirrer speed was set at lOOOrpm, sufficient to suspend all particles. Pumping of the leach slurry was conducted using peristaltic pumps and three types of Masterflex tubing, C-Flex, Teflon and polycarbonate. 3.4.2 Temperature control To achieve a leach temperature of 85°C or 95°C a Haake C l water bath was used to heat and pump water to the water jacket of each leach vessel. To inhibit evaporation of water from the water bath an insulated l id was used in addition to float balls and slow refilling of the bath, at a low flow rate so as not to decrease the bath temperature. The water bath took about two hours to heat prior to the start of the experiment with the aid of additional heaters, which were removed once the experiment had started. A copper pipe delivered the hot water from the water bath pump outlet to reinforced plastic tubing, wrapped in insulation tape, which was used to circulate the water from water bath to L S 1 , from L S I to LS2 and then from LS2 back to the water bath. Hose clamps were used to tightly secure connections and a water circulation test was performed a day in advance of each experiment to check for leaks. Teflon coated thermocouples, connected to an external digital reader, were inserted via screw fittings through the lids of both L S I and LS2 to monitor temperature. Readings were noted at the beginning, middle and end of the residence time for each leach cycle. 3.4.3 Solid-liquid separation The original intention for the countercurrent circuit was to incorporate two mini-thickeners for S-L separation for L S I and LS2 slurries. This was hoped to achieve a more continuous process and reduce the errors and difficulties previously associated with using pressure filters for all S-L 40 separation. However as this mini-thickener was a new design and had not been used before, it has was only used for L S I S-L separation in this work to investigate function capabilities. 3.4.3.1 Mini-thickener for LSI slurry S-L separation A glass mini-thickener (see Figures 14 and 15) 20cm in length, featuring a domed l id (with an internal feed well and four available inlet/outlet ports), a cylindrical glass separating column and a cone shaped underflow exit, was designed for bench-scale S-L separation. The glass l id and body were manufactured by Canadian Scientific Glassblowing Ltd, in Vancouver. A slow moving internal titanium rake, featured in Figure 16, was designed and manufactured via Falconbridge and delivered with an overhead motor, to operate slowly at 6 rpm to aid moving the settled solids through the bottom exit. A support-stand for the mini-thickener was made in the U B C Department of Metals and Materials Engineering ( M M A T ) workshop. The support consisted of a curved aluminum backing plate attached to a retort stand and large (15cm) diameter hose clamps were used to hold the thickener to the stand. Clamps were used to hold the l id in position. A thick, clear, half inch diameter piece of P V C tubing 10cm long was attached to the end of the underflow port which was then connected to smaller diameter tubing before reaching a peristaltic pump. This tube was pinched closed with clamps whilst solids settled. 41 Figure 14 Mini-thickener design Figure 15 Photo of mini-thickener The original design of the thickener included side overflow ports to assist in achieving continuous operation (i.e. slurry continually fed to the thickener would result in clarified solution exiting via the overflows). These side ports were sealed and not used in the current work due to bench scale limitations, largely with respect to liquor volumes, and instead a semi-continuous type operation was performed. This was generally achieved by pumping slurry into the central feed well , via an inlet port in the l id, and after the solids had settled fully the clarified liquor (the P L S ) was removed by pumping liquor from above the mudline. Once the P L S was removed the underflow exit tube was undamped and underflow slurry pumped out the bottom exit. 3.4.3.2 Pressure filtration for LS2 slurry S-L separation A 1.5L capacity pressure filter, manufactured by Advantec M F S Incorporated and lined with • polyurethane, was used to filter solids using N 2 at 30-45psi. It was found that for filtering finer solids two filter papers were required to prevent blowout and loss of solids. 3.4.4 Compressed air/oxygen addition to LS2 For goethite model experiments air or oxygen was required for leach stage two to regenerate the oxidant species. The gas ran from a cylinder through clear P V C tubing via a flow meter and then bubbled through a stoppered flask of water before reaching LS2 , to signal blockages. The air tube inside the leach vessel was a 9.5mm (3/8 inch) diameter polycarbonate tube about 21cm long. 43 3.5 Experimental procedures The original idea for incorporating the mini-thickener in the experimental set-up was to achieve continuous S-L separation. Whilst the mini-thickener did enable an inline, smooth countercurrent transfer of LS1 slurry, into partially leached solids and pregnant leach solution (PLS), the actual experiments in this thesis were performed semi-continuously. This means that within each experiment a series of leach cycles were performed as batch operations in continuous succession. This was largely due to limitations of bench scale work, i.e. with respect to volumes o f solution and space, and due to this being the first ever commissioning o f the mini-thickener, upon which further tests can build. A photo of the set-up is shown in Figure 17. Figure 17 Photo of countercurrent leach set-up 44 The continuous countercurrent leach experiments began with 135g of feed concentrate and 1L of feed liquor in both L S I and LS2 . A t the end of the designated residence time countercurrent transfer would occur, new feed concentrate was added to L S I and fresh feed liquor added to LS2 , and this leach/transfer pattern would repeat for 5 or 6 cycles. The first three cycles allow the countercurrent system to reach steady state operation and the following cycles were expected to approach "steady state" and thus indicate true extraction behavior. Further details of the procedures for the leach experiments follow in three sections, 1) prior to start-up, 2) during the leach cycle and 3) S-L separation and countercurrent transfer. 1) Prior to start-up The feed liquor was prepared at least a day in advance, allowing time for the solution to cool. The redox potential and p H of the feed solution were then measured at room temperature. The water bath generally took 2hrs to heat (with additional heaters) and circulate water at the operating temperature before the experiment could start. During this time, 1L of feed liquor was added to both L S I and LS2 . 2) During the leach cycle Once the leach temperature had stabilized, redox potential and temperature were recorded in each leach vessel and then concentrate was added. Temperature and redox potential were then recorded at the midpoint and end of each leach cycle. The p H was measured only in samples taken from each leach stage once they had cooled to room temperature on the bench. During the leach the following were monitored: waterbath operation, temperature and water level; stirrer connections; thickener rake motor off between transfers, and underflow clamp secured before next loading of the mini-thickener; water condenser water flowing; and hose clamps secured on water jacket tubing connections. About 15 min prior to transferring L S I slurry to the thickener, 10-15mL of 1000 ppm Superfloc® A110 anionic flocculant was added to enable quick and complete settling of solids. 45 This approximated a dosage of 100-150g per ton of leach solids. The higher dosage (15mL A110) was found necessary for experiments employing finer particle sizes. 3) S-L separation and countercurrent transfer A t the end of the leach residence time, all o f the LS1 slurry was pumped from LS1 to the glass-thickener. A t the same time, all o f LS2 slurry was pumped to a flask for immediate pressure filtration, separating final LS2 leach residues (for further washing and drying) and liquor for transfer to L S I . The LS2 filtrate was measured and recorded prior to addition to L S I , whilst allowing solids in the thickener to completely settle. Figure 18 shows the countercurrent transfer operation and volumes generally transferred. Figure 18 Mass and volume transfers 135g feed cone. LSI ~1.1L PLS -0.8L Thickener ~1.1L to 1L feed liquor L ^ l L S 2 - 1 . 2 L Pressure Filter U/F -0.2L 100% LSI solids s.d. 40-50%' Final Residue [* = 80mL sample] Pre-heated fresh feed liquor, 1L, was then added to LS2 and, following complete settling of thickener solids (see Figure 19), the clarified P L S was pumped from above the solids bed, generally leaving around 180-200mL of liquor with 100% of solids for underflow transfer to LS2 . After the P L S had been removed and measured, the underflow was pumped to LS2 . This process required the operation of the slow moving internal rake to effectively move all settled solids through the underflow bottom exit and prevent any solids clinging to walls or cone of the glass thickener. Reverse pumping was sometimes used to recollect solids that were left behind. Total transfer time took between 20-30min. Slightly before the underflow transfer to L S 2 was complete, fresh concentrate (135g) was added to L S I and this signaled the beginning of the next leach cycle. The temperature of both L S 1 and L S 2 at the start of the next leach stage was 46 recorded, and generally this was slightly lower (-5-1OC) than the designated leach temperature. However, within 5-10min the leach temperature had been reached. Figure 19 Thickening operation 47 A n example o f the experimental logs used to record temperature, redox potential and p H measurements, samples, and mass and volume transfers can be found in Appendix 6. 3.6 pH and redox potential measurements Due to problems of probe fouling in previous chloride leach experiments performed at U B C , probes were not left in the leach vessels for any length of time. Hot chloride solutions are very aggressive and combined with solids the resulting slurries are very harsh environments for glass and platinum sensors. Care was taken with both p H and redox electrodes to keep them clean between readings. When refillable probes were used the hole was covered with tape, to prevent slurry or solution entering the inside chamber of the probe. 3.6.1 pH Due to the difficulty of preventing p H fouling in hot concentrated chloride solutions, it was determined best to take p H measurements only at room temperature in the filtered samples. Measurements of p H were initially taken with plastic gel filled p H probes that had a paper junction for equilibrium ion transfer. However, the high concentrations of iron in the leach solutions quickly fouled these junctions, turning them brown, and performance rapidly deteriorated. For example, calibration was very slow and soon drift and shortening of the range made calibration nearly impossible. A new probe was then purchased, a ThermoOrion refillable combination electrode with a glass-body, rugged bulb (protected with a plastic shell), and ceramic junction. Cleaning of this probe was significantly, better, via disposing of the contaminated inner solution and soaking of the ceramic junction. Refil l ing of the electrode with 4 M KC1 (saturated with A g C l ) restored the inside environment and maintained electrode operation. The p H measurements quoted in the results section have not been corrected for liquid junction potential, largely due to the complexity of the concentrated chloride leach solution. Therefore, p H readings mainly serve to indicate comparative trends between the experiments and possibly the values at which copper or iron precipitation may occur. Readings were also generally at ~ p H 0 or lower, significantly under the lowest p H buffer of 1. Free acid measurements in the feed and final P L S provide a better indication of acid consumption and behaviour. 48 3.6.2 Redox potential Redox potential, E , measurements were originally performed using a glass body combination A g / A g C l electrode with a ceramic junction and fine platinum sensor. Not only did the fine platinum sensor bend in the hot leach slurry, after only a few trial experiments the glass body cracked without warning simply upon insertion of the probe into the slurry. This happened twice and for one experiment a Saturated Calomel Electrode was thus used as an immediate backup. A n epoxy body A g / A g C l combination electrode ( V W R Symphony Brand) with a flat bottom platinum sensor and free flow junction was subsequently purchased. The filling solution used was 4 M KC1 saturated with A g / A g C l , suitable for use in concentrated solutions (>0.2M). Light 's solution (see Appendix 5) was used to check redox electrode performance. Redox potential readings were taken in-situ in the leach vessels at the start of the experiment, before addition of concentrate, and at the middle and end of each residence time. During the leach Light's solution was used to check performance of the electrode, and whether drifting or fouling of the electrode had occurred. Generally readings were satisfactory and only sometimes the KC1 filling solution would need renewing. Redox potential measurements were not corrected for temperature and have been quoted in this thesis as m V vs A g / A g C l at either room or experimental temperature. The measurements have also not been corrected for Light's solution changes during the experiments, unless in one or two cases where they were out by greater than 30 m V . E readings can therefore generally be accepted as accurate to within 20-30mV. The leach solutions in this research work are highly concentrated and depart significantly from ideal solution behaviour. Therefore, the redox measurements largely serve a comparative purpose between experiments in this thesis. 3.7 Sampling and Analysis 3.7.1 Sample Collection L S I and L S 2 80mL slurry samples were collected at the end of each residence time, during transfers to the thickener and pressure filter. Approximately 200mL was first allowed to flush 49 through the pump line and then 80mL was collected in a beaker. The samples were then set-aside on the bench for filtration. Both L S 1 and L S 2 samples were prepared in the following manner. The sample was filtered via water vacuum filtration and the filtrate was collected in a small (60mL) plastic sample bottle. The sample was cooled to room temperature, prior to dilution for external analysis and p H measurement. The solids were then washed with lOx diluted concentrated hydrochloric acid (1/10 HCI) until the filtrate ran colorless. This chloride wash solution was collected for all residue washes, and the total volume was recorded arid sent for external analysis. Final washing of the filter cake was with DI water and the cake was then transferred to an oven to dry. Once dry, the weight of the residues were recorded and sample bagged for external analysis. The final LS2 leach residue sample, collected in the pressure filter, was also treated in the same manner as the residue in the slurry samples. For this residue the weight of the wet cake was recorded in addition to the washed and dried product. A l l solution and residue samples were prepared accordingly and sent to LPL for analysis. The final L S 2 residue was sent to ChemMet for sulfur group analysis. 3.7.2 Preparation and Analysis The following samples were sent to IPL for analysis. The feed liquor and L S I and LS2 liquor samples from each cycle were diluted lOOx (2mL in 200mL) with -2 .5% H C I and analyzed for 30 elements (including C u and Fe) by ICP. In addition, undiluted samples of the feed liquor and final exit P L S were sent for free acid analysis and ICP 30 element analysis. Dilutions with only DI water for the feed (lOOx dilution) and final exit P L S (lOOOx dilution) were also prepared for total chloride ion analysis via ion selective electrode (ISE). 50 A l l solids were sent for 30 element ICP analysis (via aquaregia digestion). The feed concentrate was also analyzed for Total S and Si02. Two further samples were sent to ChemMet Consultants, the feed concentrate was analyzed for A u and the final L S 2 exit residue analyzed for sulfur group analysis (including total sulfur, elemental sulfur, sulfide and sulfate). 3.7.3 Calculation of copper and iron extraction A n example calculation for determining the extraction of copper and iron can be found in Appendix 7. 51 4 RESULTS AND DISCUSSION 4.1 Excel mass balance outputs for the hematite and goethite process flowsheets This section reports the input and output streams of the mass balance for the two chloride process models. For both models 135g pure CuFeS2 concentrate was introduced to the circuit and it was assumed that 100% of the C u and Fe in CuFeS2 was leached and subsequently all of the extracted C u was removed via S X and all extracted Fe was precipitated (i.e. steady state was maintained), either as goethite or hematite. The models were simplified to include only major reactions, assumed from previous experimental work and the literature. These reactions are reported and discussed. Also , the simplified models consider complete countercurrent transfer of liquor with respect to solids. The models do not account for the small portion of L S I liquor that would return with thickener underflow to LS2 . For each stream, the copper and iron species are written as single, uncomplexed cations (i.e. C u + , C u 2 + , F e 2 + , Fe 3 + ) but are assumed to be associated with chloride. These copper and iron species would generally be added to the leach liquor as metal chloride salts and thus contribute to the total background chloride concentration. A concentration of 110 g/L CaCL. was used to ensure a concentrated chloride medium >5M. In these simplified models, gypsum (CaS04) precipitation is not considered, as minimal sulfate production is expected. A background concentration of HC1 was also included in both models although remained constant through the balancing of acid consuming and generating reactions. The result of ensuring this acid balance also assisted in regenerating the oxidant species for recycling to Leach Stage 2. 52 4.1.1 The hematite process This model focuses on 3 major process units (see Figure 20), a 2 stage countercurrent leach, copper solvent extraction and iron precipitation in an autoclave. Thus separating the leach residue from iron deportment and producing two waste residues. The input and output streams for L S I , LS2 , S X and Autoclave, determined from the mass balance for this flowsheet, are shown below alongside the assumed reactions. The mass of each species is given in grams (g) and the total amount of water, lOOOg, is comparable to the quantity used in the experiments and is maintained constant, assuming evaporation of water produced. Figure 20 Hematite Process Feed Concentrate, 100% CuFeS2 135g CuFeS 2 (added to L S I ) : C u = 46.7g; Fe = 41. lg ; S = 47.2g Assumed extent of extractions: L S I = 51.5 %; LS2 = 48.5 % Leach Stage 2 (LS2) CuFeS 2 + 4 F e C l 3 -» 5FeCl 2 + C u C l 2 + 2S° (4.1) Liquor, rich in ferric and cupric, is recycled from the autoclave and fed to LS2 . Ferric is consumed in the oxidation of CuFeS 2 , leaving some ferric and cupric for further leaching in L S I . Table 9 Hematite Process: L S 2 streams Stream C u + C u 2 + F e 2 + F e J + C a C l 2 H C I H 2 0 In 0 50.3 0 80.2 110 2 1000 Out 0 73.0 99.6 0.53 110 2 1000 53 Leach Stage 1 (LSI) CuFeS 2 + 4 F e C l 3 -> 5 F e C l 2 + C u C l 2 + 2S° (4.2) CuFeS 2 + 3 C u C l 2 -» 4 C u C l + F e C l 2 + 2S° (4.3) Ferric is completely consumed in L S I while the remaining CuFeS 2 is leached with cupric. Sulfur oxidation and atmospheric oxidation of cuprous or ferrous is not considered in this simplified model. A s the feed is 100% CuFeS 2 the consumption of acid by soluble minerals is not considered. The exit stream is predominantly reduced species, i.e. cuprous and ferrous. Table 10 Hematite Process: LSI streams Stream C u + C u 2 + Fe Fe C a C l 2 HC1 H 2 0 In 0 73.0 99.6 0.53 110 2 1000 Out 95.5 1.54 121.3 0 110 2 1000 Solvent Extraction (SX) 2 C u C l + 0.5O 2 + 2 H R o r g C u R 2 o r g + C u C l 2 + H 2 0 (4.4) Addition of oxygen to the aqueous phase facilitates 100% extraction of the leached copper (i.e. - 5 0 % of the copper in solution) into the organic phase. To balance acid generated, from the donation of protons from the organic extraction, an equal amount of cuprous is oxidized to cupric. Table 11 Hematite Process: Solvent Extraction streams Stream C u + C u 2 + F e 2 + F e J + C a C l 2 HC1 H 2 0 In 95.5 1.54 121.3 0 110 2 1000 Out 2.01 48.3 0 80.2 110 2 1000 54 Hematite Precipitation in Autoclave (A/C) 2 F e C l 2 + 0.5O 2 + 2 H 2 0 -» Fe 2 0 3 (s) + 4HC1 (4.5) 2 C u C l + 0.5O 2 + 2HC1 -> 2 C u C l 2 + H 2 0 (4.6) 2 F e C l 2 + 0.5O 2 + 2HC1 -> 2 F e C l 3 + H 2 0 (4.7) Oxygen is added to the autoclave to enable the precipitation of 100% of leached iron as hematite (Fe 2 0 3 ) via ferrous oxidation. This reaction produces a surplus of acid that is subsequently consumed by cuprous and further ferrous oxidation. This acid balance also achieves 100% regeneration of the required ferric and cupric for recycle to LS2 . Table 12 Hematite Process: Autoclave streams Stream C u + C u 2 + F e 2 + F e J + C a C l 2 H C I H 2 0 In 2.01 48.3 121.3 0 110 2 1000 Out 0 50.3 0 80.2 110 2 1000 4.1.2 The goethite process This process focuses only on 2 major process units (see Figure 21). Though the leach is more complex than the hematite model, as oxygen is added to precipitate goethite in LS2 , and thus a mixed sulfur/iron residue is produced. The input and output streams for L S I , L S 2 and S X , determined from the mass balance for this flowsheet are shown below alongside the reactions considered. A s above with the hematite model, the mass of each species is given in grams (g) and the total amount of water, lOOOg, is comparable to the quantity used in the experiments and is maintained constant, assuming evaporation of water produced. Oxygen addition is not included below but is assumed to be equal to that required. Figure 21 Goethite Process 55 Feed Concentrate, 100% CuFeS2 135g CuFeS 2 (added to L S I ) : C u = 46.7g; Fe = 41 . lg ; S - 47.2g Assumed extractions: L S I = 55 %; LS2 = 45 %. Leach Stage 2 (LS2) CuFeS 2 + 3 C u C l 2 -» 4 C u C l + F e C l 2 + 2S° (4.8) 2 F e C l 2 + 0.5O 2 + 2HC1 ^ 2 F e C l 3 + H 2 0 (4.9) 2 F e C l 3 + 4 H 2 0 -» 2 F e O O H ( s ) + 6HC1 (4.10) Overall - 2 F e C l 2 + 0.5O 2 + 3 H 2 0 -» 2 F e O O H ( s ) + 4HC1 (4.11) Cupric in solution, recycled from solvent extraction, facilitates leaching of CuFeS 2 . Oxygen is added to enable oxidation of ferrous and subsequent precipitation as goethite. Wi th the excess acid generated from goethite precipitation, cuprous and further ferrous oxidation occurs. The oxygen input required is 17.65g. Table 13 Goethite Process: LS2 streams Stream C u + C u 2 + Fe F e J + C a C l 2 H C I H 2 0 In 0 48.7 31.1 0 110 2 1000 Out 0 69.7 0.27 8.21 110 2 1000 Leach Stage 1 (LSI) CuFeS 2 + 4 F e C l 3 -» 5FeCl 2 + C u C l 2 + 2S° (4.12) CuFeS 2 + 3 C u C l 2 -> 4 C u C l + F e C l 2 + 2S° (4.13) The small amount of ferric generated in LS2 is consumed in CuFeS 2 leaching and then the remainder of the leach is conducted using cupric. The pregnant leach solution is rich in copper, as cuprous, and the iron content is the same as fed to LS2 , as there are no further downstream reactions involving iron. 56 Table 14 Goethite Process: LSI streams Stream C u + C u 2 + F e 2 + F e i + C a C l 2 HC1 H 2 0 In 0 69.7 0.27 8.21 110 2 1000 Out 93.5 1.93 31.1 0 110 2 1000 Solvent Extraction (SX) 2 C u C l + 0.5O 2 + 2 H R o r g -> C u R 2 o r g + C u C l 2 + H 2 0 (4.14) This step is the same as for the previous hematite model. V i a the addition of oxygen, cuprous is oxidized to enable extraction into the organic phase and complete cupric regeneration, for recycle to LS2 . Table 15 Goethite Process: Solvent extraction streams Stream C u + C u 2 + F e 2 + - F e 3 + C a C l 2 HC1 H 2 0 In 93.5 1.93 31.1 0 110 2 1000 Out 0 48.7 31.1 0 110 2 1000 4.1.3 Feed liquor outputs For both process models, the concentrations of copper and iron species in the recycled L S 2 feed stream, as calculated from the above mass balance, are summarized in the following table. These were used for the preparation of feed liquor for the countercurrent leach experiments. Table 16 Feed liquor concentrations Model Feed Requirements Hematite F e J + 80 g/L, C u 2 + 50g/L, HC1 ~3g/L Goethite F e J + 0 g/L, F e 2 + 30 g/L, C u 2 + 50g/L, HC1 L o w (<3g/L) 57 For the hematite model, sufficient ferric and cupric must be fed to LS2 to enable 100% extraction of C u and Fe from CuFeS2, as there is no additional oxidant addition during L S I or LS2 . For the goethite model, ferric is not required in the L S 2 feed as oxygen is added to produce ferric for goethite precipitation and regenerate cupric for further leaching. One aspect of these models is that ferric is consumed in leaching prior to cupric. Whilst this may be true for a medium with a total chloride concentration of - 4 . 5 M , where the standard F e 3 + / 2 + potential is at 668mV and C u 2 + / + at 581mV, at higher chloride concentrations it is predicted that the copper couple w i l l become more oxidizing than the iron couple. Wi th respect to the models, in which 100% of CuFeS2 occurs and there is no excess oxidant in the feed, it does not matter for the output feed which oxidant is consumed first. For example, in the hematite model both oxidants are completely consumed, and in the goethite model only a small amount of excess ferric is produced, which is subsequently consumed in CuFeS2 leaching. 4.2 Overview of continuous countercurrent leach experiments Twelve continuous countercurrent leach experiments have been performed. The order, experimental identifier ( ' H ' or ' G ' ) and conditions used are shown in Table 17. The experimental identifier indicates upon which process model the leach was based, ' H ' representing the hematite (no air in leach) process and ' G ' the goethite process (with air/oxygen addition to LS2) . The residence time (RT) quoted in the table is per each of the two leach stages. The first eleven experiments were conducted using Antamina concentrate, varying conditions of residence time, temperature and particle size. The final experiment was performed using fine grind Rosario concentrate, at optimum conditions determined from the eleven previous experiments. For all experiments, to start each experiment 135g of feed concentrate was added to volumes of 1L feed solution, and the background concentration of total chloride, contributed by 110 or 220 g/L C a C ^ , was aimed at greater than >5M. The aim of these experiments was to find conditions to achieve maximum copper extraction (>95%) by varying conditions of residence time, temperature and particle size, with both leach models to be investigated. The first four experiments were important in establishing experimental procedures and providing direction for subsequent experiments. 58 Table 17 Order of experiments Order Expt Feed Concentrate RT (h) Total Time (h) Temp. '(Q Air/0 2 to LS2 (flow rate) 1 H I Antamina (p90=75um) 2 10.75 85 N o 2 G l 2 10 85 A i r (200mL/min) 3 H2 2 10 95 N o 4 G2 2 10 95 A i r (2.5L/min) 5 H3 3 18 85 N o 6 H4 r 3 18 95 N o 7 H5 Antamina (p90=41um) 3 18 95 N o 8 H6 2 12 95 No 9 H7 3 18 85 N o 10 H8 2 12 85 N o 11 G3 r 3 18 95 0 2 (2.5L/min) 12 H9 Rosario (p90=37um) 3 17.5 95 N o For the first four experiments 5 leach cycles were performed, but for subsequent experiments 6 cycles were performed. The first three cycles are necessary for the system to reach steady state, and then actual trend data can be collected. Thus it was considered better to have three end cycles rather than two to provide the trend data. Some anomalies in total experiment duration exist for two experiments, H I and H9, due to problems with the waterbath. In both cases the total leach time was extended, to make up for leach time when temperatures were depressed, but overall did not significantly affect the output of the experiments. This is discussed further in the following results sections. Also in the first experiment, H I , redox measurements were taken using an S C E probe due to a glass body probe breaking immediately before start-up. For all subsequent experiments an epoxy body A g / A g C l probe was acquired and proved to better withstand leach conditions. For the first 59 two experiments, H I and G l , p H measurements were measured on the bench at temperatures slightly above room temperature. However for the remainder of experiments p H was consistently measured only at room temperature. For the goethite experiments, G l and G2, oxygen addition was required and this was achieved via sparging compressed air to LS2 , in accordance with previous U B C chloride leach experiments. A similar flow rate to those previous experiments, of 200mL/min, was used for experiment G l but upon recalculation, was later found to be inadequate. It was determined then that 535mL/min would deliver the exact amount of air required (containing - 2 1 % oxygen) over two hours and so a 4.7x excess should be used in experiment G2 to overcome inefficiencies, i.e. 2.5L/min. The impact of the greater airflow delivery to the small leach vessel resulted in a depression in leach temperature of - 4 C in L S 2 to around 91C. This excess air, however, still did not achieve a significantly greater extraction and so the experimental focus shifted to the hematite leach model for the next 6 experiments. For flow rate calculations see Appendix 7. The next 2 experiments, H3 and H4, were also performed using the coarse Antamina concentrate completing a series of four hematite model experiments comparing residence time (2h vs 3h) and temperature (85 vs 95C). Then four experiments, H5-H8, again based on the hematite (no air in leach model) were performed using Antamina concentrate that had undergone fine grinding. These four reflected the experimental matrix of the four previous coarse material experiments (i.e. 2 or 3 h R T at 85 or 95 C) . The final two experiments were performed at optimum conditions determined for the hematite model experiments, using fine grind concentrate, a temperature of 95C and residence time of 3h, applied to the goethite model and an alternate concentrate. This included one further goethite model experiment (G3) but this time oxygen (98% pure) at a flow rate of 2.5L/min (an excess of 33x) was sparged to leach stage two. The final experiment, H9, was conducted based on the hematite (no air in leach) model at optimum conditions (3h, 95C) using finely ground Rosario concentrate. 60 The following sections in this chapter present the results of the countercurrent leach experiments. Certificates of analysis for all samples can be found in Appendices 1 and 2, and the Cu and Fe mass accountability performed for all experiments, based on feed material in and samples taken out, can be found in Appendix 8. Profiles of measurements over time for each individual experiment, can be found in the Appendix 9. 61 4.3 Results of hematite model experiments Eight experiments based on the hematite leach model were conducted and conditions are shown in Table 18. The first four experiments employed the as received Antamina concentrate (p90 = 75pm), varying residence time (2 or 3 hour per leach stage) and temperature (85 or 95C). The successive four (H5-H8) employed fine grind Antamina concentrate (p90 = 41pm), similarly varying residence time and temperature. Thus an experimental matrix was completed observing the effect of particle size, residence time and temperature on copper extraction. Table 18 Hematite model - feed conditions Experiment (RT, Temp.) Concentrate C l [M] E m V X Ag/Ag/CI (a = vs SCE) C u 2 + g/L Fe 3 + g/L Free Acid (as HCI) g/L H I (2h, 85C) Antamina (p90=75um) 7.19 626a -0.48 47.8 76.7 5.74 H2 (2h, 95C) Antamina (p90=75um) 7.39 650a -0.24 47.2 79.9 5.28 H3 (3h, 85C) Antamina (p90=75um) 7.19 639a -0.36 52.9 87.2 4.04 H4 (3h, 95C) Antamina (p90=75um) 7.95 661 -0.35 55.8 90.6 4.04 H5 (3h, 95C) Antamina (p90=41um) 6.29 677 -0.16 45.0 76.0 2.71 H6 (2h, 95C) Antamina (p90=41um) 6.35 660 -0.25 41.1 70.2 1.94 H7 (3h, 85C) Antamina (p90=41um) 8.01 659 -0.68 47.7 80.1 2.96 H8 (2h, 85C) Antamina (p90=41um) 7.17 661 -0.67 48.0 75.3 3.69 "measured at room temperature After the first four experiments it was observed that the best conditions for copper extraction involved using a 3h R T at 95C. Therefore for the fifth experiment these conditions were used when first testing the fine grind concentrate in order to see potentially what the maximum extraction would be and i f further experiments with this model were worthwhile. Upon achieving very good extractions, >98%, the final three experiments were pursued. 62 The feed liquor cupric and ferric concentrations, with respect to the mass balance output, were aimed to be 50g/L and 80g/L respectively. In a few of the experiments these concentrations were slightly lower, however as the feed concentrate was not 100% CuFeS2 (unlike in the model mass balance calculations) this decrease of oxidant in the feed liquor was not a concern. A recalculation of oxidant in the feed (see Appendix 7), based on the actual CuFeS2 content of the Antamina feed shows that 70g/L ferric and 40g/L cupric are sufficient oxidant concentrations for the feed. Temperature was maintained within 1-2C from the target leach temperature, except in experiments H I and H3 when there were slight deviations (due to problems with the water bath). In H I the temperature in the second cycle dropped, thus the leach was extended for 45 min to compensate, and at the end L S 2 was slightly elevated at 88C for the last hour. In H3 there was a slight drop of 2-4C in the last cycle for a period of ~20min, due to a water jacket hose clamp connection loosening. These deviations are not considered significant. The copper and iron mass balance was calculated for each experiment based on the total in feed materials and outgoing samples, solutions and residues. A l l of the balances were within acceptable limits, of 95-105%) accountability, except for copper in H I , for which copper accountability was 94%, and H3 , for which.copper accountability was 93%. A s these were only out by 2% they were not considered significant. 4.3.1 Potential, pH and free acid A l l 8 hematite model experiments, involving Antamina concentrate, exhibited similar potential and p H profiles and trends in free acid consumption. The input feed and final output LS1 and L S 2 streams are summarized in Table 19. The E and p H profiles for each experiment are found in Appendix 9 and some representative profiles have been included below. In accordance with the trend predicted for countercurrent leaching, potential decreased from L S 2 to L S I . L S 2 is expectedly higher in the countercurrent system due to the fresh feed liquor meeting the partially leached solids from L S I . In these experiments there was a general redox potential gradient of 50-100mV from LS2 down to L S I , with in-situ redox measurements for LS2 at 450-550mV A g /A g ci and L S I at 400-450mV A g /A g ci • Examples of representative E profiles for experiments H4 and H5 are shown below in Figures 22 and 23. 63 Figure 22 E profile for H4(75um, 3h, 95C) Figure 23 E profile for H5(4lum, 3h, 95C) The pH profiles for all hematite model experiments revealed a similar trend, showing only slight deviations from the feed value. For all hematite model experiments the free acid was measured in the exit PLS stream and found to be completely consumed. The pH therefore was not affected by the decrease in free acid, owing to the extremely high concentrations of chloride and metals. The free acid was likely consumed in the oxidation of reduced species. 64 Table 19 E, pH and free acid output for hematite experiments Expt Stream E mV Ag/Ag/ciX (except a = vs SCE) pH (at room temp, except b = at 35C) Free Acid as HC1 (g/L) H I Feed 677 a -0.48 5.74 L S I 416 a -0.01 b not measured LS2 509 a -0.04 b H2 Feed 733 -0.24 5.28 L S I 453 -0.39 0.18 LS2 530 -0.41 H3 Feed 750 -0.36 4.04 L S I 451 -0.34 O . 0 1 LS2 530 -0.31 H4 Feed 747 -0.35 4.04 L S I 440 -0.55 O . 0 1 L S 2 537 -0.37 H5 Feed 754 -0.16 2.71 L S I 415 -0.27 O . 0 1 L S 2 492 -0.13 H6 Feed 735 -0.25 1.94 L S I 428 -0.42 <0.01 LS2 497 -0.33 H7 Feed 748 -0.68 2.96 L S I 388 -0.77 <0.01 LS2 468 -0.71 H8 Feed 733 -0.67 3.69 L S I 400 -0.41 O . 0 1 LS2 462 -0.73 "At reaction temperature 65 4.3.2 Copper and iron extraction The following Table 20 summarizes solids and solution concentrations in the feed and final LS1 and L S 2 output streams. The calculated copper and iron extraction values in the final L S I and L S 2 residues are also shown. The profiles of solution and residue concentrations for each experiment can be found in Appendix 9 and representative profiles have been included in the following discussion of the results. These experiments revealed a consistent trend with respect to CuFeS2 leaching in varying temperature and residence time with respect to particle size. The copper extraction at steady state in the first four experiments (H1-H4), employing p90 = 75um concentrate, show that increasing the temperature from 85C to 95C had a greater impact on extraction than increasing the residence time from 2h to 3h. Observing the final LS2 extraction data, for experiments H l -H4, the copper extractions were 80% (2h,85C), 87% (2h,95C), 83% (3h,85C)and 93% (3h,95C) respectively. A s >95% copper extraction in chloride media has been well demonstrated by others, the subsequent experiments (H5-H8) were performed using a concentrate of a finer particle size, p90 = 41 um, in order to achieve higher copper extraction. The first experiment in this series, H5 , was thus performed using conditions of 3h residence time and 95 C. This experiment was very successful and an average of ~ 98% copper extraction was achieved. The final three experiments, H6-H8, were then completed to mirror the experimental matrix performed for the p90 = 75um Antamina concentrate, i.e. varying residence time (2 and 3h) and temperature (85 and 95C). The lowest average copper extraction was 90% when a 2h leach stage residence time was employed and conducted at 85C, see H8. Interestingly, experiments H6 (2h, 95C) and H7 (85C), both reported similar copper extraction behavior and a final value of 96%. The lower values reported for iron extraction have been examined and found to represent iron remaining from unleached CuFeS2 and iron remaining as undisturbed pyrite (FeS2). Examples of representative profiles for solution concentration, C u and Fe in L S 2 residue and C u and Fe extraction for experiments H4 and H5 are shown in Figures 26-31. 66 Table 20 Results of hematite model experiments Expt Stream Solution Solids Percentage Extn Cug/L Feg/L Mass g Cu % Fe% Cu% Fe% HI Feed 47.8 76.7 135 28.8 28.4 LSI 74.1 101.7 87.6 18.0 23.0 59 47 LS2 54.1 85.2 44.5 17.5 22.1 80 74 H2 Feed 47.2 79.9 135 28.8 28.4 LSI 81.8 114.3 79.3 17.9 19.6 64 59 LS2 60.1 93.1 63 8.1 12.4 87 80 H3 Feed 52.9 87.2 135 28.8 28.4 LSI 86.3 125.1 88.9 20.7 22.0 53 49 LS2 63.0 100.7 57.7 11.3 16.2 83 76 H4 Feed 55.8 90.6 135 28.8 28.4 LSI 94.6 127.9 46.6 21.4 24.3 74 70 LS2 68.1 102.9 46.9 6.1 12.5 93 85 H5 Feed 45.0 76.0 135 28.3 28.8 LSI 86.0 118.0 114 25.0 24.7 27 28 LS2 71.0 100.0 41 2.0 8.5 98 91 H6 Feed 41.1 70.2 135 28.5 28.7 LSI 75.3 103.9 100.1 25.2 24.8 34 36 LS2 60.8 87.4 46.6 3.7 9.6 96 88 H7 Feed 47.7 80.1 135 28.5 28.7 LSI 81.1 112.9 71.5 20.7 21.8 61 60 LS2 65.4 96.9 43 4.0 10.2 96 89 H8 Feed 48.0 75.3 135 28.5 28.7 LSI 80.4 106.5 87 23.0 24.9 48 44 LS2 63.8 91.5 49.1 7.9 13.36 90 83 67 Leach Conditions: H4 = 75p.m, 3h, 95C and H5 = 41ujn, 3h, 95C Figure 26 H4: Cu & Fe (%) Extraction Figure 27 H5: Cu and Fe (%) Extraction 93.4 94.0 91.2 9 2 9 5 10 Time (h) 15 20 Figure 28 H4: Concentration vs Time Figure 29 H5: Concentration vs Time 140 120 100 ! 80 60 40 20 0 -4—LSlFe - • — LS2 Fe - • — L S l C u « LS2Cu 5 10 Time(h) 15 20 140 Figure 30 H4: Cu and Fe in LS2 Residue Figure 31 H5: Cu and Fe in LS2 Residue 68 4.3.3 Speciation Both models predict that the exit P L S w i l l contain predominantly reduced species, o f iron and copper, and hence downstream purification and recovery steps were balanced accordingly. Theoretical calculations (based on the actual feed concentrate, feed liquor and completeness of leaching) for total reduced species (TRsp) content in the P L S have been performed (an example calculation is included in Appendix 7) and presented in Table 21 along with actual titration values. Both results are compared and variations from the model are discussed, as are implications for downstream processing. Table 21 Total reduced species of exit PLS and L S 2 liquors Expt CuFeS2 in (mol) Final Cu extn % CuFeS2 reacted (mol) TRsp (theoretical) LSI (M) Ttl theo. M TRsp (titration) LSI (M) Ttl actual M TRsp Difference (Theoretical -Titration) (M) Titration TRsp % of Theoretical TRsp (titration) L S 2 (M) H2 0.612 87 0.532 2.66 3.24 2.44 3.33 0.22 92 1.02 H3 0.612 83 0.508 2.54 3.41 2.28 3.60 0.26 90 0.78 H4 0.612 93 0.569 2.84 3.64 2.60 3.78 0.24 91 0.88 H5 0.605 98 0.593 2.97 3.26 2.60 3.47 0.37 88 0.88 H6 0.605 96 0.581 2.91 3.07 2.53 3.05 0.38 87 1.48 H7 0.605 96 0.581 2.91 3.35 2.62 3.30 0.29 90 1.22 H8 0.605 90 0.545 2.72 3.19 2.39 3.17 0.33 88 1.30 The calculation for determining the theoretical reduced species concentration takes into account the feed liquor oxidant concentrations and the actual CuFeS2 reacted for each experiment. It does not matter which oxidant, ferric or cupric, is consumed first in leaching, as both reactions produce the equivalent amount of reduced species, i.e. 5 mol reduced species per mol CuFeS2 reacted, as shown in equations 4.15 and 4.16. CuFeS 2 + 4Fe 3 + -> 5Fe 2 + + C u 2 + + 2S° Reduced = 5 mol ferrous (4.15) CuFeS 2 + 3Cu 2 + 4Cu++ F e 2 + + 2S° Reduced = 4 mol cuprous and 1 mol ferrous (4.16) 69 The titration values reported in the above table are all lower than the theoretical calculation by 0.22-0.33M, and in both cases oxidant was still found remaining in the P L S . To explain the difference between theoretical and titration total reduced species values two things were considered, additional oxidation occurring in the leach liquor during the experiment and error in the titration method. The first considers that additional cuprous or ferrous may be oxidized by atmospheric oxygen, available in the headspace of the leach vessel, with available acid, HC1. However a calculation of air volume and acid required indicate that between 7-12L of air would be the exact amount of air required to deliver the required oxygen and 8-12g/L HC1 would be required. Whilst there may certainly be acid available in the leach solution it is not feasible that the solution would meet with such a volume of ambient air, considering that there is only 1L of headspace in both leach stages. Also an excess o f air, above the stoichiometric ratio, would be required to overcome gas transfer limitations. Although not all of the difference can be attributed to atmospheric oxidation a small amount is certainly possible. It is more likely that cuprous, being incredibly unstable in more dilute solutions of chloride on exposure to air, may have been oxidized during dilution for the titration. In either case, both results indicate that copper and iron species in the P L S stream are not likely to be fully reduced as predicted in the models and downstream implications must be considered. Complete reduction of the liquor could be achieved by adding copper metal to the stream to reduce the oxidized species prior to downstream processing. This however would not only increase the cost of production but also increase the copper tenor, which is already high. Another alternative is to investigate what the actual impact would be of sending this liquor containing oxidized copper as well as reduced species direct to solvent extraction. The result is actually not detrimental as the following discussion suggests. The uptake of copper by the organic extractant is actually in the form C u 2 + and thus oxidized species would be available without requiring addition of oxygen. However without the complete balance of oxidation of cuprous to cupric there would be an excess of acid produced. Though ferrous is still available and may be oxidized thus consuming acid produced by copper extraction. The following equations (4.17-4.19) demonstrate these reactions. 70 (4.17) (4.18) (4.19) Thus the stream leaving S X would contain cupric, ferrous and some ferric, which is then available for hematite precipitation. 2 F e C l 3 + 3 H 2 0 -» Fe 2 0 3 (s) + 6HC1 (4.20) as opposed to 2 F e C l 2 + 0.5O 2 + 2 H 2 0 F e 2 0 3 ( s ) + 4HC1 (4.21) This may produce a slight excess of acid in the autoclave. 4.3.4 Hematite experiment discussion Conditions for achieving the highest extraction of copper involved employing a fine grind concentrate, 95 C and leaching for 3h in each leach stage. Depending on the copper recovery necessary there may be some flexibility with respect to leach design, due to fairly high extraction rates reported at the lower temperatures or leach residence times. For example, i f >95% recovery is sufficient, then using fine grind concentrate and either a 2h residence time at 95C or a 3hr residence time 85C may be sufficient. Should fine grinding of the concentrate be avoided then it may be possible to conduct the leach at 95 C with a 3h leach residence time and still achieve a satisfactory copper extraction of around 93%. 4.4 Results of goethite model experiments Three experiments based on the goethite leach model have been performed and conditions are shown in Table 22. In the first two experiments air (-21% 0 2 ) was added to leach stage 2 to regenerate cupric for further leaching and ferric for goethite precipitation. These first two experiments were not successful in achieving high copper extractions and so for the third experiment fine grind Antamina concentrate was tested and 98% oxygen used as the oxidant source. The flow rate for the third experiment was kept the same as for the second, i.e. 2.5L/min. This was to ensure that an excess (circa. 3 Ox) of oxygen was delivered to overcome inefficiencies and solubility limitations. The calculations for oxygen delivery are contained in Appendix 7. 71 C u 2 + + 2 H R o r g C u R 2 o r g + 2 H + 2 C u C l + 0.5O 2 + 2HC1 -» 2 C u C l 2 + H 2 0 2 F e C l 2 + 0.5O 2 + 2HC1 -» 2 F e C l 3 + H 2 0 Table 22 Feed conditions for goethite model experiments Exper iment 3 Flow rate to L S 2 C l [M] E m V X Ag/Ag/Cl p H C u 2 + g/L Fe 3 + g/L Fe 2 + g/L C a C l 2 g/L Free Acid (as HC1) g/L G l (2h, 85C) Air (200mL/min) 4.5 463 0.02 54.1 7.9 27.1 110 0.72 G2 (2h, 95 C) Air (2.5L/min) 5.7 457 -0.47 55.9 11.6 22.9 220 10.02 G3 (3h, 95C) 0 2 (2.5L/min) 5.6 443 -0.29 47.4 4.9 25.1 220 <0.01 (residence time per leach stage, temperature); measured at room temperature In accordance with the calculated feed requirements, liquor concentrations approximated 50g/L cupric, with 30g/L iron as ferrous and a low concentration of ferric. Although experiment G3 had a slightly lower cupric concentration, this was not important as there was additional oxidant available in the feed as ferric and the feed was not 100% CuFeS2, unlike the mass balance model. In experiment G l the background concentration of HOg/L CaC^ provided just under 5M total chloride, as a result of the lower content of metal chlorides used in the feed compared to the hematite feed. Thus for further goethite model experiments the background CaCL. concentration was raised to 220g/L to ensure a total chloride concentration of 5M. A high chloride content is important for supporting high metal concentrations and maintaining a low pH, however this increase was not thought to be significant for comparison of these experiments. It was intended that the feed liquor contain a low (<3g/L) concentration of free acid as HC1 to promote goethite precipitation. The second experiment G2 however was prepared with a significantly higher HC1 content. Whilst this was not a concern regarding copper extraction it was interesting to observe the effect on iron extraction behavior. The temperatures for the experiments were kept within 1-2C of the set conditions, except for LS2 in experiments G2 and G3 where the temperature was depressed ~4C due to the high flow rate (2.5L/min) of air and oxygen addition. 72 The copper and iron mass balance was calculated for each experiment based on the total feed liquor and concentrate in and outgoing samples, solutions and residues. A l l o f the balances were within acceptable limits, o f 95-105% accountability, except for experiment G l in which C u accountability was 88% and iron accountability was 86%. One problem that may have contributed to this poor recovery was the difficulty processing and filtering the sticky, orange goethite residue. Another problem with this experiment was that insufficient oxygen was delivered to LS2 and so overall this experiment can only serve to provide an approximate guide for goethite model leaching. The results of the above goethite model experiments are reported below. 4.4.1 Potential, pH and free acid The measurements of redox potential, p H and free acid for the input feed and final output L S I and LS2 streams for the three goethite experiments are summarized in Table 23. Similarly, in accordance with the trend predicted by countercurrent leaching, potential decreased from L S 2 to L S I . A s the potential of the feed liquor was quite low and the air (-21% 0 2 ) addition was not very high in the first experiment the E gradient from L S 2 to L S I was not very large. In experiment G2 when air addition was substantially increased, the resulting potentials of L S I and L S 2 were still not observed to increase and did not differ from G l . The resulting extraction results, as discussed below, were not very high either and this is further discussed below. Thus air delivery, even in excess, did not seem successful at regenerating oxidant species or even restoring potential to that of the feed value. A significantly higher oxidation potential was however achieved in the final experiment, G3 , with L S 2 between 550-600mVAg/A gci, when almost pure oxygen (98%) was added to L S 2 at a rate of 2.5L/min and thus delivering a 20 times excess. A substantial E gradient was observed across L S 2 / L S 1 . This E increase also paralleled a greater copper extraction. Figures 32 and 33 show the E profiles for experiments G2 and G3. This experiment confirmed that copper extraction could be improved by increasing oxidation potential in LS2 and it is likely that >95% copper leach extraction could be achieved with longer leach residence time and further fine grinding of the concentrate. 73 Table 23 E, pH and free acid output for goethite experiments Expt Stream E mV Ag/Ag/ciX (LS1/LS2 for feed) PH Free Acid as HCI (g/L) G l Feed 504/524 0.02 0.72 LSI 439 0.75y 1.5 LS2 451 0.64y G2 Feed 534/571 -0.47 10.02 LSI 439 -0.04 2.8 LS2 456 0.54 G3 Feed 515/580 -0.29 <0.01 LSI 409 -0.24 <0.01 LS2 606 0.32 x At reaction temperature;y measured at 50C. Figure 32 E profile for G2(75|im, 2h, 85) Figure 33 E profile for G3(41u.m, 3h, 95C) 580 560 540 5 520 I 500 < <n 480 |> 460 | LU 440 420 400 • LS1 • LS2 * • * • • 4 6 8 Time (h) 10 12 74 Fairly similar p H profiles were observed for all goethite experiments, as shown in Figures 34 and 35. A s expected there was greater fluctuation in p H for the goethite experiments, compared to the hematite model, due to L S 2 reactions involving acid, i.e. ferrous and cuprous oxidation and iron precipitation reactions. During the first leach cycle, prior to steady state, the p H in LS2 rose sharply but then returned to stabilize at a lower p H . This reflects acid consumption reactions dominating, i.e. oxidant regeneration, until a sufficient acid deficit can drive goethite precipitation. This occurs significantly in G3, in which the feed liquor already had a lower free acid content. In G3 , the p H was observed to rise sharply to a value of 2, which is the value at which iron is expected to precipitate in concentrated chloride solutions. Following this first cycle a better balance was achieved between oxidation and precipitation reactions stabilizing the pH at a lower value. A slightly different situation occurred in experiment G2 due to the feed liquor containing a greater concentration of free acid. A s a result it appears that p H is stabilized at a moderately low value, and it is likely that oxidation regeneration dominated due to the available concentration of free acid in the feed, and only a small amount goethite precipitation could occur when sufficient acid had been consumed. Figure 34 pH profile for G2(75u,m, 2h, 95C) Figure 35 pH profile for G3(4lu,m, 3h, 95C) 75 4.4.2 Copper and iron extraction The following table summarizes solids and solution concentrations in the feed and final L S I and L S 2 output streams. The calculated copper and iron extraction values in the final L S I and L S 2 residues are also shown. The profiles of solution and residue concentrations for each experiment can be found in Appendix 9. The results of sulfur analyses are given in section 4.6. Table 24 Output concentrations (solution and solids) for goethite model experiments Expt Stream Solution Solids Percentage Extn C u g / L Feg /L Mass g C u % F e % C u % F e % G l Feed 54.1 35.0 135 28.8 28.4 L S I 55.5 27.6 90 25.5 25.4 41 40 L S 2 67.3 29.6 112 13.9 29.8 60 13 G2 Feed 55.9 34.5 135 28.8 28.4 L S I 87.9 59.7 93 29.1 26.9 30 35 L S 2 76.7 48.5 78 16.6 23.2 67 53 G3 Feed 47.4 30.0 135 28.5 28.7 L S I 93.9 26.2 88 24.5 26.5 44 40 LS2 79.3 11.3 108 4.0 38.4 89 -7 Experiments G l and G2, at 2h R T and 85C and 95C respectively, were able to demonstrate leaching only to 60% and 67% copper extraction. Therefore in these conditions, despite delivering a 4.7 times excess of oxygen in experiment G2, copper extraction was not acceptable. Oxygen solubility in hot liquor at atmospheric pressure is l ikely to be the limiting factor. Wi th respect to iron, precipitaion of goethite in G l successfully recovered all but 13% of iron, driven by a low free acid in the feed liquor. However, in G2, iron precipitation was inhibited by the presence of lOg/L HC1 in the feed liquor, although this had no apparent consequence on copper extraction. The increase of 8% copper extraction from experiment G l to G2 mirrors the 7%) increase in the hematite experiments when a 2h leach stage residence time was used at 85C 76 the 95C, i.e. H I and H3 . The extraction profiles for copper and iron in experiments G l and G2 are shown in Figures 36 and 37. The best result for the goethite model was experiment G3 using Antamina fine grind concentrate (3h R T , 95C) and oxygen (98%) sparging to LS2, instead of air. Fol lowing the first three cycles, copper extraction stabilized at 89% and a much higher oxidation potential in LS2 was recorded. The extraction profiles for copper and iron in experiment G3 is shown in Figure 38. A longer residence time and/or finer grinding may lead to achieving >95% copper extraction. Interestingly however iron precipitation was too high in G3 , i.e. greater than the amount o f iron contained in the feed concentrate, resulting in a negative -7% iron extraction. A greater amount of acid in the feed would reduce iron precipitation, however the value of free acid would be fixed by solvent extraction step for this model. The current precipitation value may however be acceptable providing a greater overall extraction of Cu and Fe from CuFeS2 can be achieved, or by ensuring that further cycles through the system w i l l achieve an overall iron balance. 77 Figure 38 G 3 : C u & Fe (%) Extract ion -80 J — — Time (h) 4.4.3 Speciation The following table presents the titration values for total reduced species concentrations, along with total species determined from the addition of copper and iron concentrations, for the feed and final L S I and LS2 streams. Similarly as with the hematite model, the exit P L S streams all contain a significant portion of oxidized species. However, as the next process step in the goethite model flowsheet is copper solvent extraction the impact on downstream process is insignificant, as discussed previously for the hematite model experiments. 78 Table 25 Total reduced species in goethite experiments Expt Stream Total Species (M) Total reduced species (M) % Total Reduced species G l L S I 1.37 0.65 48 L S 2 1.59 0.55 35 G2 L S I 2.45 1.41 58 L S 2 2.08 0.94 45 G3 L S I 1.95 1.12 58 L S 2 1.45 0.09 6 4.4.4 Goethite experiment discussion The goethite experiments were not successful in achieving greater than 95% copper extraction. Even when conditions were used that had achieved the highest extraction in the hematite model, i.e. fine grind concentrate, 3 hour leach and 95C, and an excess of oxygen was delivered to the leach the highest extraction achieved was 89%. This value was a significant improvement on the experiments employing air addition to LS2 , G l and G2, and by extending residence time and/or finer grinding of the concentrate there is potential for achieving >95% copper extraction. Although such a large excess of oxygen (circa. x20), as used in experiment G3, would be very unreasonable in scaling up of the process. It is l ikely that there exists an industrial scale reactor designed to increase oxygen utilization efficiency in a commercial plant, thus improving upon the bench scale system. Another negative aspect of this model was that the mixed, orange sulfur/goethite leach residue was very sticky and difficult to filter. Not only was filtering slow but the residue encapsulated a significant amount of liquor which required large amounts of 1/10 H C I for washing away soluble copper and iron. 4.5 Application of hematite leach model to Rosario concentrate The highest copper extraction from the Antamina concentrate in this investigation was achieved employing the hematite model with a fine grind concentrate (p90 = 41pm), leach residence time 79 of 3h and a temperature of 95C. The aim of the final experiment, H9 , was to achieve a similarly high copper extraction from a different concentrate under the same conditions. The Rosario concentrate was selected as it is currently of interest. The Rosario concentrate contains a similar amount of copper (26.8%) as the Antamina (28.5%), but contains less CuFeS2 and an additional copper mineral, chalcocite, CU2S. The Rosario concentrate underwent fine grinding prior to the experiment, to a p90 of 37p.m. The conditions of the feed liquor for the Rosario experiment are shown in the following table, and are similar to the conditions for the previous hematite model experiments, HI -8 . Table 26 Feed conditions for Rosario experiment C u 2 + Fe 3 + Free Acid Exper iment 3 C l E m V C a C l 2 [M] p H g/L g/L g/L (as HC1) Ag/Ag/Cl* g/L H9 7.08 659 -0.41 50.6 82.0 110 2.35 (37um,3h,95C) (residence time per leach stage, temperature); measured at room temperature During cycles 5 and 6 of this experiment water bath circulation problems significantly disrupted the temperature profile. Interestingly, copper extraction, as discussed below, was not significantly affected. The mass balance accountabilities for copper and iron, determined from feed and outgoing samples and materials, were 95% and 97% respectively. 4.5.1. Potential, pH and free acid The potential, p H and free acid trends were similar to those of the other hematite experiments and the results are shown in Table 27 and Figures 38 and 40. The free acid was similarly depleted in the P L S , the p H stayed close to the feed value and the range o f L S I and L S 2 E values were 400-450mVAg/A gci and 450-550mVAg/A gci respectively. 80 Table 27 E, pH and free acid Expt Stream E mV Ag/Ag/ciX pH Free Acid as HCl(g/L) H9 (37um,3h,95C) Feed 743 -0.41 2.35 LSI 463 -0.51 <0.01 LS2 562 -0.53 x At reaction temperature Figure 39 E profde for H9(37um,3h,95C) Figure 40 pH profile for H9(37um,3h,95C) 800 750 700 650 600 550 500 450 400 350 —•—LSI I — • — LS2 -4 6 8 Time (h) 10 12 14 4.5.2 Copper and iron extraction Table 28 reports the successful output data for the final hematite experiment. Not only was the final copper extraction 98%, by the end of the first leach stage 96% extraction had already been achieved. The presence o f CU2S in this material thus indicates that it is easier to leach than a predominantly CuFeS2 concentrate, and, as a result thereof, the required leaching time is significantly decreased. The following equation (4.22) is shown for CU2S leaching. The unleached iron in the residue was examined and found to contain represent undisturbed pyrite. Cu 2 S + 2Fe 3 + /Cu 2 + / ox -> 2Fe 2 + /Cu + / red- + 2 C u 2 + + S° (4.22) 81 Table 28 Output concentrations (solution and solids) for the Rosar io experiment Exp t Stream Solution Solids Percentage E x t n C u g / L F e g / L Mass g C u % F e % C u % F e % H9 Feed 50.6 82.0 135 26.8 26.8 (37LUIL L S I 91.7 100.5 76.3 2.1 28.3 96 40 3h, 95C) LS2 55.1 84.9 68.6 1.1 27.5 98 48 Figure 41 C u & Fe (%) Extract ion for H9(37Lim,3h,95C) 0 5 10 15 Tlme(h) 82 4.6 Sulfur balance The feed concentrate was analyzed for Total sulfur (S), and the final L S 2 leach residue of each 2 2 experiment was analyzed for total S, elemental S, sulfide (S ") and sulfate (SO4 ")• The sulfur group analyses in the final LS2 residue of each experiment are shown in Table 29. Table 29 Sulfur species in Final LS2 residues Expt Elemental (%) Sulfide (%) Sulfate (%) H I 28.0 23.9 0.91 H2 50.4 15.0 <0.1 H3 39.6 16.6 <0.1 H 4 42.0 13.9 <0.1 H5 49.8 9.91 <0.01 H6 56.5 11.8 O . 0 1 H7 53.6 11.4 0.02 H8 46.8 14.4 O . 0 1 H9 23.1 34.1 O . 0 1 G l 15.0 16.0 1.33 G2 24.2 21.3 0.12 G3 20.4 6.68 1.09 The residue (<10g) from the slurry sample taken from L S I was not however analyzed for sulfur, and to calculate the sulfur mass balance an expected loss of sulfur from L S I was determined based on C u and Fe concentrations in the sample. Table 30 reports these sulfur values and presents the sulfur balance. The generally very low amounts of sulfate indicate almost 100% yield of elemental sulfur. The hematite model experiments produced a mainly elemental S residue, with some sulfide remaining from unleached CuFeS2 and/or FeS2. For the goethite model experiments, elemental sulfur was also the major reaction product left in the residue along with goethite (FeOOH). For both models, there was no significant amount of sulfate in the exit leach residues. This correlates 83 previous reports of minimal oxidation Of elemental sulfur to sulfate, which would precipitate with gypsum in these experiments. A small but insignificant amount of sulfate was detected in the first and final goethite experiments, the two in which goethite was successfully precipitated with the residue. Table 30 Sulfur recovery Expt Total S in (g) LS2 Output Mass (g) Total Sulfur in LS2 (%) S out of LS2 (g) Est'd S out of LSI sample (g) Total Sout (g) M S N out/in (%) Sin-S out (g) Final Cu ext'n (%) H I 43.3 44.5 53.2 23.7 4.0 27.7 64 15.6 80 H2 43.3 63.0 65.6 41.3 3.3 44.6 103 -1.3 87 H3 43.3 57.7 56.2 32.4 2.9 35.4 82 8.0 83 H4 43.3 46.9 55.9 26.2 2.5 28.7 66 14.6 93 H5 43.8 41.0 59.8 24.5 3.5 28.0 M 15.7 98 H6 43.8 46.6 68.3 31.8 3.4 35.2 So 8.6 96 H7 43.8 43.0 65.0 28.0 3.2 31.2 71 12.6 96 H8 43.8 49.1 61.1 30.0 3.7 33.7 MBSM 10.1 90 H9 48.0 68.6 57.2 39.2 6.1 45.3 95 2.6 98 G l 43.3 112.4 32.6 36.6 2.8 39.4 VI 3.9 60 G2 43.3 77.8 45.5 35.4 2.4 37.8 87 5.5 68 G3 43.8 108.2 28.2 30.5 4.1 34.6 79 9.1 89 The sulfur balance for nearly all experiments does not lie within acceptable limits, 95-105%, despite acceptable copper and iron balances. There appears to be a loss of S in the range of 5-15g across all experiments. Reasons for this are not clear other than to consider the possibility of loss of residue or slurry. It was observed in the hematite experiments, that the light sulfur residue would tend to float to the top and stick to the walls of the flask or thickener. This indicates potential for loss of sulfur from the final residue and may have a significant effect on the balance. It is indeed the sulfur 84 balance of the hematite experiments that report the greatest lack of accountability of sulfur (up to 15g). The three goethite model experiments and the final Rosario leach experiment reported better sulfur recovery figures, with a lack of accountability between 2.6-9. l g . These residues contained a significant amount of goethite or pyrite that may have helped bond the residue together and prevented floating or frothing of the lighter sulfur. In scaling up of the process, attention to engineering design would address the lack of sulfur accountability in the leach residue. 4.7 Trace Metals A l l the elements reported in ICP analysis were examined to determine which other elements were leached with copper. The following table reports the extraction values of these elements in the final LS2 residues of three experiments, and includes copper and iron extractions for comparison. Silver (Ag), chromium (Cr), lead (Pb), magnesium (Mg), nickel (Ni) and zinc (Zn) all reported significantly high extractions, i.e. all generally above 90%. This indicates the suitability of chloride leaching for complex sulfides. Table 31 Trace metal analysis Element Extraction % in final L S 2 residue H5 (41um,3h,95C) G3(41um,3h,95C,O2t0LS2) H9 (Rosario,3h,95C) Ag 97.1 92.4 98.9 Cd not in feed not in feed 100.0 Cr 92.1 46.2 93.9 Cu 98.2 88.7 97.9 Fe 91.1 -7.0 47.7 Pb 99.9 99.7 94.3 M g 96.7 93.4 95.5 N i 93.0 90.1 92.4 Zn 98.9 98.4 98.7 85 5 CONCLUSIONS AND RECOMMENDATIONS 5.1 Model mass balances Two new process flowsheets, the Goethite Process and the Hematite Process, were developed and the successful completion of mass balances shows potential for both circuits particularly in addressing current purification difficulties and waste management issues. Solvent extraction (SX) is known to be a very effective method of selective uptake of copper from sulfate leach liquor, for transfer to the electrowinning circuit with minimal impurity transfer, and is expected to work well for a chloride system. A s goethite (FeOOH) is a less stable iron residue in the environment that hematite (Fe20s) it is important to have two process options because in some localities goethite w i l l not be accepted in waste for disposal. The mass balances were also useful in determining feed liquor requirements for the leach. Although these models were simplified, considering only major reactions and assumed complete extraction of chalcopyrite, they were still beneficial and successfully tested in the continuous countercurrent leach experiments. Whilst the aqueous processing assumptions were fairly accurate with respect to the experimental outcomes, the delivery of gas containing oxygen did not perform as wel l as the goethite model intended in an atmospheric leach system. Both Goethite and Hematite Processes did however improve upon a previously proposed flowsheet in which air was added to leach stage two in addition to use of an autoclave downstream for iron recovery, thus requiring careful control so as not to precipitate iron in the leach. In the two current models investigated in this thesis, either oxygen was added to the leach to precipitate iron or an autoclave was incorporated downstream, therefore clearly defining at which part of the process iron deportment would occur. 5.2 Implementation of the mini-thickener The mini-thickener proved very useful in achieving consistent operation for all countercurrent continuous-batch experiments. Complete transfer of the solids was achieved from L S I to L S 2 and the transfer of leach liquor volumes could be carefully tracked. In previous U B C chloride leach testing, a more continuous countercurrent leach system was investigated in which half the slurry was removed and countercurrently transferred at each half-residence time. However, it was difficult to accurately take out half each time and use of the pressure filter, for S-L separation, was difficult to peform cleanly. The transfer time in this investigation was still 86 comparable to previous work, at 20-30min, but as the L S I S-L separation was achieved inline via the mini-thickener the transfer operation was much smoother. Following from this initial testing of the mini-thickener, it is feasible that future work could potentially incorporate two mini-thickeners, for both L S I and LS2 S-L separation, in addition to working towards a more continuous operation. This would likely involve slurry being continuously fed to the thickener and the overflow ports being used to slowly remove clarified pregnant leach solution. 5.3 Countercurrent leach experiments The countercurrent leach experiments performed for the hematite model were successful in demonstrating the effect of varying particle size, leach time and temperature to achieve copper extractions >95%. There are three hematite process variations possible i f 96% extraction is sufficient. A l l three involved fine grind concentrate (p90=41pm) with either a 2h residence time at 95C (96% extraction), a 3h residence time at 85C (96% extraction) or a 3h residence time at 96C (98% extraction). The benefits of this model include a relatively straightforward 2-stage leach (in which all the oxidant required is fed as aqueous species to LS2) and the separation of sulfur leach residue from iron precipitate, thus making reuse or waste treatment easier. The substitution of N a C l for CaCL; may be possible for the hematite model as there is no iron precipitation in the leach, and hence there is no concern for jarosite formation, as in previous Falconbridge Copper Chloride Process experiments undertaken at U B C . This substitution would reduce processing cost. The goethite model experiments were unsuccessful in achieving >95% extraction of copper, with the highest extractions stabilizing at 89%. However a longer residence time or finer grinding may be investigated to achieve higher extractions for this model. Alternatively, use of an autoclave for Leach Stage 2 may overcome oxygen solubility limitations and in turn may assist in the increase of copper extraction. Wi th respect to fine grinding, although potentially costly and energy intensive, grinding to 20microns is considerably more feasible than grinding to lOmicrons or less, as is required for some competing sulfate based leach processes. 5.4 Recommendations for further work Following the success of the leaching aspect of the hematite model it would now be recommended to test the subsequent process units with respect to the actual exit liquor arising from the leach or relevant previous process unit. Thus for solvent extraction a feed should be 87 tested containing variations of reduced and oxidized copper and iron species, instead of a assuming a feed containing only reduced species. Further work should also involve development of a more continuous countercurrent leach employing the use of two mini-thickeners for L S I and L S 2 S-L separation. 88 R E F E R E N C E S 1. Peters, E. 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(1987) The leaching of chalcopyrite with cupric chloride, Metallurgical Transactions B, V0I.I8B: 31-39. 33. O'Malley and Liddell, K . C . (1987) Leaching of CuFeS 2 by aqueous FeCl 3 , HC1 and NaCl: Effects of solution composition and limited oxidant, Metallurgical Transactions B, V0I.I8B: 505-510. 90 34. Dutrizac, J.E. (1990) Elemental sulphur formation during the ferric chloride leaching of chalcopyrite, Hydrometallurgy, 23:153-176 35. Lotens, J.P. and Wesker, E. (1987) The behaviour of sulphur in the oxidative leaching of sulphidic minerals, Hydrometallurgy, 18:39-54 36. Chen, T.T. and Cabri, L.J . (1986) Mineralogical overview of iron control in hydrometallurgical processing, Chapter 1:19-55 in Iron control in hydrometallurgy, Eds: J.E. Dutrizac and A.J . Monhemius, Ellis Horwood series in industrial metals. Chichester : E. Horwood ; New York : Halsted Press. 37. Riveros, P.A. and J.E. Dutrizac (1997) The precipitation of hematite from ferric chloride media, Hydrometallurgy (46) 85-104. 38. Hoffman, J.E. (1991) Winning copper via chloride chemistry - an elusive technology, JOM, August, 48-49. 39. Hyvarinen, O. and Hamalainen, M . (1999) Method for producing copper in hydrometallurgical process, US Patent, 6,007,600. 40. Schweitzer, F.W. and Livingstone, R.W. (1982) Duval's C L E A R hydrometallurgical process, In: P.D.Parker (Ed), Chloride Electrometallurgy, TMS-AJME, New York, 221-227. 41. Moyes, J. and Houlis, F. (1999) The INTEC copper process, Copper '99, Arizona. 42. Moyes, J. (2000) A revolution in copper - the Intec copper process, January, Sydney, Australia. 43. Dalton, R.F., Diaz, G., Price, R. and Zunkel, A . D . (1991) The cuprex metal extraction process: recovering copper from sulfide ores, JOM, August, 51-56. 44. Ferron, C.J., Williamson, R.G. and Zunkel, A . D . (1996) Metal recovery from a complex sulphide concentrate using a ferric chloride leach process, Minerals, EPD Congress, California, Metals and Materials Society/AME (USA),181-192 . ' 45. Demarthe, J .M., Gandon, L. and Georgeaux, A . (1976) A new hydrometallurgical process for copper, In: J.C.Yannopoulos and J.C.Agarwal (Editors), Extractive Metallurgy of Copper, T M S -A I M E , New York, 825-848. 46. Demarthe, J .M. and Georgeaux, A . (1978) Hydrometallurgical treatment of complex suphides, Complex Metallurgy '78, IMM,113-120 47. Satchell, D.P. and Gerlach, J.N. (1986) Chloride hydrometallurgical process for production of copper, US Patent, 4,594,132. 48. Fletcher, A.W. , Sudderth and Olafson, S.M. (1991) Combining sulfate electrowinning with chloride leaching, JOM, August, 57-59. 49. Kordosky, Gary (1999) Copper solvent extraction coupled with concentrate leaching, a short course presented at Copper '99, Arizona. 91 50. Kordosky, Gary (1999) Solvent extraction and electrowinning of copper from concentrate leaching solutions I, solvent extraction of metals,: a short course presented at Copper '99, Arizona. 51. Tinkler, O. (1999) Solvent extraction and electrowinning of copper from concentrate leaching solutions II, hydrometallurgical processing of concentrates, Copper '99, Arizona. 92 Sample identification guide for Appendices 1 and 2 The following table provides a guide indicating to which experiment the samples in the certificates of analysis belong, e.g. H I has an original experiment identification number of 7 and ' M ' and ' N ' represent L S I or LS2 samples respectively. The concentrate samples analyzed are also identified in Table 33. Table 32 Sample identification guide Experiment Identifier Original Identification Number Samples L S 1 . L S 2 1 H I 7 7 M , 7 N 2 G l 8 8P, 8Q 3 H2 9 9K, 9L 4 G2 10 10J, 10K 5 H3 11 11C, 11D 6 H4 12 12M, 12N 7 H5 13 13G, 13H 8 H6 14 14V, 14W 9 H7 15 15S, 15T !0 H8 16 16X, 16Y 11 G3 17 17K, 17L 12 H9 18 18 A , 18B Table 33 Concentrate sample identification Concentrate Sample I.D. 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" S £ to n ™ -ts Z g S i j C 0 c = o -a • S c w O f l j a j 9- E ' - C M c O ' ^ - m c o c D T - c M c O ' ^ - i n c o S S E ' c T ^ ^ s E _, u . u - X X X X X X X > - > - > - > - > - > - 5 < m O O £ Z Q O CO 3 © ( O t O C O < O C O t O C O C O C O C O C O ( D ( D < D C O C O ( O C O © E CO z x: c D^-CMCOM-LOCDCDI-CMOOTJ-IOCD^ »c ™ 3©(O (0COtOU}COCO<OtOtOCDCOCO<0COCDCDCO X) C O i - C O C f t O ^ - f O C O t O l O l O CL ^ CM IO (VI i f I O O) O ( O CL o COC CMCMCMCMCMCMCMCMCMCMCMCMCMCNCM CN O Q_ C C f S m W N f ' I ' O O O ' - ' T ' l f N i - * O Q_ —I o. V V V V V V V V V V V V V V V O o N § _ K i r O L D ^ ' - L n r O C O ^ M N r t C O N O O f j o — y- cocnmcMcnr^cocMcococMcococMco o = IIIII1I1111ISI1SP 5i llllllllllllll! Is o | 1 8 i s 8 i i l B 5 i | 8 ! § i p 3 | 2 3 3 3 3 S 2 3 3 2 2 2 2 S S 2 | & m E CL C L CMCMCMCMCMCMCMCMCMCMCMCMCMCMCM V V V V V V V V V V V V V V V I O Q_ CM 3 s 5 l l l l l l l l l l l l l l s §§f g „ I I I I I I I I I I I I I IS P § j P l ffSf££Sffffffff i p E Q. CL • ' - t N ' - ' - n i - ' - i - i - O CL 8 9 3 | g S | K S S 3 S ? S ^ S g S 8|fe £I V V v v v v v V V V V V V V V ^Sfe * | s § s s § 2 S s s s s s a s g §p ' i Iff!!!!!!!!!!!! §j& l O ) N r l f i O t ' " W ' - i - r ( l ) N n N t o - r - r j r t V r - i - v T - CM o o c o i a j T f w t D i n ^ i f t ^ i f l c n i o m c o u ) »- o i 5 K C M s i o o o n n i n » - t ' 9 , o > n o ) ' < r o n C M T - T - T - C M ^ C M C M C M C O C O C M C M ^ - C O C < < < < < < < < < < < < < < » - • r - O ' K Ol C C O O N N ' t f C B O J N i n c O i - C M S C n O O O Q. O ° ^ ^ ^ ^ ^ ^ ^ ^ ^ ^ i ^ ^ ^ ^ C O O O ^  S o . K C M C O O l O C O C M C O O C O C O C O r ^ O J L O O O o Z Z Z Z Z Z Z Z Z Z Z Z Z Z ^ b *~ ? CL o j r M O J > - r - j f > j ( N f \ j ( N j i - i - » - T - r - T - T - o r co CM I g 0 - C L Q . Q . a a C L a . C L Q . C L C L Q . Q . C L Q. 1 1 0. 0. Q. Q. Q. Q. H Q. Q. Q. Q. Q. CM CN g 5J B E a a co ) C 0 C 0 C 0 C 0 C 0 C 0 t 0 C 0 C 0 C 0 C 0 C 0 C 0 ( •CMCO^-incOCOT-CMCOCO^lOCO( I I ' Q. WWCOCOCOCOCOtOCOCOCOCOCOCOC .o_i mm<D'ff(D'to)ioco5so)v*t!0< p o ^ ^ T ^ i o N l i D i d r i n o d tri c v co CM m co_i ininiAcniniA(Dino)0)(piriininu)(o< O O O O O O O O O O O O O O O O S c d d d d d d d d d d d d d d d d 2  vv vv v v v v v <D_1 i n C D C O I ^ O O C O O ^ ^ C p O l i - C M C O C M C O ^ 0)iDNU5ioo^(D(b o tn cn co to C M T - ^ T - C M C M C M C M L O T - T- O cn CM v-CM ^ ' ^ d ^ c s J C M ^ c r j ^ i o d u S o r ^ ' f O t D ^ S ( D O N r M A n ( J t O r ( O J I I O O ) T -^ 0 ) ( 0 0 0 ) 0 ) 0 > 0 ) C O O > N e O S N S V m T- T - cn cn O o) o t - o * - o . - * - ' - c o » - o o o o d 0 0 = s E d ^ d d d d d d i D d d d d d d 2 p O N O d d d d i o d d d d d d ® 2 C CM V co_j c>jiAco(nTrsNSinrM(oin(otororo< BNininminmifin in in m m 3 5 T3_i T-T-^T-fmininnm--rt'-'-'-s< O "5, o o o 0 o o o o o o o o o o o o * - ^ E o d d d d d d ^ d d d d d d d 2 gj sJ '-i-coc>inmioi->-in<tiDCBC>ii-T-< o) d o d T-" ^ d d d d d d d d d d ^ co-j T - c n ^ ^ - T - c o ^ - ' - ' - c o c M C M C N i C M ^ - c n r f CO ^  p c o p p p p p p c o p p o o o o o E s E d d d d d d d d d d d d d d d d 2 «J,_J CMCMCMCMCMCMCMCMlOCMCMCMCMCMCMCM< cn d d d d d d d d i N d d d d d d d ? E V V V V V V V V V V V V V V V m cn O CD o g in cn o cn o 0 1 co cn o co d g> cn o cn 6 g> *- cn o cn d g> cn o cn d g> T - cn d g CO i- cn o cn d 0 1 CO d g CO i- cn o co d 0 1 CM cn d g co N-gj O W O O O O O O Q O O O O O O O ^ E d d d d d d d d d d d d d d d 2 c _ i (O'-nioinnNNMONCiiini-moirf N "9) ^ q ^ o o N O J t p c n c v i i o ^ c q v a q ^ c d d v c S ' T f f l i r i i ' i N d N r s l f f l s r i ' c V * - C O C N C M C M C N C 0 C M CO CM "~ ^ O ^ O O O O O O C O O O O O O O O ^ E d ^ d d d d d d r d d d d d d d z cn O O O O O O O O O O O O O O O O ^ E V V V V V V V V V V V V V V V V - 1 O) O O O O O O O O C O O O O O O O O ^ E V V V V V V V v v v v v v "c-P j-J CMCMCMCMCMCNCMCNCMCMCMCMCM o> d d d d d d d d d d d d d d d d ? E ' V V V V V V V V V V V V V V V V ^ -(Jj-J mcomcMininmmr^tnmm(Dmeo«-< E d i N d d d d d d r d d d d d d ^ 2 ra_J O K C M h - v i n ^ c o ^ i n ^ e o ^ i n i o ^ - r f ? o m -3- Z p co co D) _J C M C M i f C M ( O I O O ) i - l O S N T - T - * C O ( 0 < <e> oqT-inNcowco^cNioi-Q'Ooio^ c 0 0 0 0 0 0 0 0 ^ 0 0 0 d d c V V CM o_i cMnt- fT-^T-^m»- i - i - i -nT-o)< CO j^, OCMOOOOOOinOOOOOOO^ c d d d d d d d d d d d d d d d d ^ c V V V V V V V D> - CO - - CD ^ - CM E z OJ O O O O O O O O O O O O O O O O ^ E V V V V V V V V V V V V V V V V ^ 1 = ? ? ? = s i i 0.1 9999 ICPH20 2 l 0.06 7.47 0.25 0.69 0.32 0.51 0.44 0.5 28.31 0.24 0.12 0.18 0.12 0.15 0.18 4.47 N/A N/A 1 1 E 5.5 20.1 6.4 7.8 7.3 6.9 6.6 142.2 5.8 5.8 0.2 9999 ICPH20 i t 1 = 1 I | 5 1 1 5 | | | | | | 1 I I l i l l l l l l l l l l l i [ 1 1 1 ! ? l mmmnwm II I f l l l l l l i l l i i i i i i i l l s l l i t 0.94 57.68 1.16 1.63 1.37 1.44 1.41 1.44 114.86 1.27 1.05 1.16 1.11 1.11 1.15 11.43 N/A N/A i i ¥ I f l l l l l l f l l l l i i i ! i l l 1 | | *t s s s s s ^ - g s s s s s s - i i 5 I o CQ ID O LU , - -'C Q co o CM » u. 2 e o (p CD LMJ O CM » 2 o ™ O Ifl 00 OJ _ CL n O O O O O O O O O O O O O O O O O I- I l l l i l l l l l l l l l l l l l to o o o o o o o o o o o o o o o o o o CO tO«tf)tOWtOtOCOCOCOCOCOCOCOCOCOCOCO r - C C C C C C C C C C C C C C C C C 5 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 I l l l l l l l H I I I I I I I i o o o o o o o o o o o o o o o o o o COCOtOCOCOCOtOtOCOCOCOCOCOCOCOCOCOtO «2 2 co ; £ iJ O * - O •= T3 => < a C ID O g. j , y- • T -CMP ) * m ( D l f l^Nn* i n ( D ^ E co Z •0 Q> -o ^ O "n. 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E ^-^-D -Q.Q.Q.Q. a. d. a. a. a. a. ci. a. a. a CO Q.Q.Q.(LQ.Q.a.a.lLlQ.£llO.Q.[LCL 33 l ls ls III I 1* S 5 i i i 111 E a a s 3 : c/J a CL 5 5 t O m Hi i- „ _ c c ^ ra LU § S * ® 1 . i i isle's . i i 11 t 2 i J < s s 3 3 S | 5 a s s a s s s Iff 1 s £ £ a a I s s m 2 J $ taooadodoSoooioiaiooio 3^2 CJOCLZQQ to S !2 2 £ t : fc 1 2 J £1 —i i f l O ) i - [ M O ) t n M n c M i n m m i o i f l < D < a~r3) 0 ( 0 ( o n t o ( J c o ( D O ) t o o o o o i n < E d w d ^ d d d d m d d d d d d d 2 to—) $ i n i A i n m i o m i n u ) i A i n m m m m m < { —i ~g, 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 = 5 c d d d d d d d d d d d d d d d d z c V V V V V V V V V V V V V V V c o o i n i n o s o x o N O M D i o t o i ' - ' ' CO i - ft O O O O O O i ' - l O O l l S O l f f l O ) co t - T - T - T - ^ c n co n _ j Q S ' - t o w t o c M C M n S ' - c j ' - ' - c o s r f O o i o n ^ S c M T ^ « w ( 0 0 ) C » c » f f l c q n N < e ^ d c d i ^ i o o o c N ^ c N ^ i r i c N ^ c o ^ a i 2 i n f t £ S d ) 0 ) m a i \ i c o i n ( D u > < D U } 0 ) CO i - CM CM M- CO 0 % p co q 3 p d d q w ^ 6 d o o P ^ 5 E o m o o o O CD O O O o o o ,\ -J C N ^ C M O C M C O C D ^ f C O C P C O C O C O i n c O C O < U " g , LO CM o n i D w n o o s o s o j N f f i ^ C O CM C O C M C M C M C M O C N O ' ^ o d d l O 2 C CO T-CM CD I 0 ) N ( D ^ « l ( O I D ! O W 1 0 ' - a ) I O V ' - P ) < 0"c3> r ^ S J r o c o c N i r i d ^ c o ^ o i o i d c o ^ ^ - ' 1 ? CMSCNCNCMCMCMCMCOCMCMCMCMCMCMCO CM CM 0 cl p c o p ^ p p p p i ^ p p p q q o o ^ E d o d d d d d d c i d d d d d d d ^ T - i - f M C O « - t O ) i f i - ( O C « l t ' f m C M ' * < o d d d d d d d d d d d d d d o N T E v v v v * (0—1 i - i - M O ) ! D l p © f f l O ) ( D C ) « ' - ' - r i r < CD ^ , q Q q q q q q q i o o o o o o O ' - ^ E o d d d d d d ( \ d d d d d d d z O I N N N « « O I N ( O N C N C i l C N W N C \ l < a> d d d d d d d d i r i d d d d d d d ? E V V V V V V V V V V V V V V V .0 _) r r r r r r r r N N r r i - T - T - r < t w cb d d d d d d d d d d d d d d d d ? g v v v v v v v V V V V V V ^ < O 0 C O C M T r f ^CMCM^-CMCnCM0OCOCOCOC0< C> CMlritO COTtrOCOCDCOCOCM CM CO ^ E CM h--L ; J < < < < < < < < < < < < < < < < ^ ° g> z z z z z z z z z z z z z z z z f ^ c in CM o=d < " < < < < < < < < < < < < < < < 1 O Z p j Z Z Z Z Z Z Z Z Z Z Z Z Z Z Z < LU ^ _J < < < < < < < < £ < < < < < < < < Z Z Z Z Z Z Z Z d ^ I Z Z Z Z Z Z Z CM V r-O a. £ to CO WWWCOCOtOWCOCOCOCOCOCOCOCOCO CO o>8 i5 £ « Is-, I e « s I • S JO . -*t p l l l l l p l l l l l l l §i| *t lI13Si|||iasan § ! | > | p l p l i p i l l l f f s II| 5 | 5 5 3 5 3 5 5 2 5 2 8 5 3 5 2 5 III Ft i'iii5ii"i55i55s III • F | 2 2 8 3 3 2 2 3 3 3 8 ^ 3 3 2 2 111 * | 118 * | i p 5 l ! l l 2 i p i | P 5 I | ** PPPIPPPPI § 1 | * | O j s r . „ , g s „ „ r s | l °" | i i i i i i i i i i i i i i i i 511 2 | ipS3S3^5gillii I^I *f ISsSSSSsSsiilli* Mj ? t IHlllllllllllll | | | f | 2 S 3 ' " S 3 5 3 8 S 2 5 8 2 ! ......... i S §»gj s a s m Q w *J z fecSBflja, 2- ^ T - c M c o ^ i n c o t D T - c M c o ^ i n c D ? IK ffl cT « •£ •£ £ L L U . < < < < < < < c o m a i t D C Q C Q S < c o D = C O (5 ffl tO COCDCOCOCDCOCOCDCDCOCOCOCOeOeOCOCOCD O O 0 - Z Q D CO l - l - l - l - T - T - l - l - l - l - T - T - T - » - T - l - » - l -a> E z • D O ) TJ x; e a ^ C M c o x t w c o c D ' - C M c o ^ i o c o ^ <5 £ L L . u . < < < < < < < c o m m c t i c D c o s < c o £ <S OOCOCOOOaScOCOTOCOCOCOCOOOCOcjOCOCOCO Appendix 2 Certificates of Analysis (ChemMet) 121 Appendix 3 Certificates of Analysis (Malvern Mastersizer) 131 A S T E R SIZE R Sample Name: Ant - Whole - Average Sample Source & type: Factory Sample bulk lot ref: Result Analysis Report SOP Name: Measured by: contract Result Source: Averaged Measured: Monday, October 06, 2003 11:01:08 AM Analysed: Monday, October 06, 2003 11 01:09 AM Particle Name: Accessory Name: Analysis model: Sensitivity: Chalcopyrite Hydro 2000S (A) General "purpose Normal Particle Rl: Absorption: Size range: Obscuration: 1.450 0 0:020 1 to 2000.000 um 18.04 % Dlspersant Name: Dispersant Rl: Weighted Residual: Result Emulation: Water 1.330 0.956 % \ Off Concentration: Span : Uniformity: Result units: 0.0629 %Vol 2.267 o:72 Volume Specific Surface Area: Surface Weighted Mean D[3,2]: Vol. Weighted Mean D[4,3]: 0.246 mJ/g 24.418 um 52.146 um d{0.1): 10.613 d(0.5): 41.971 d(0.9): • 105.757 Particle Size Distribution 1 ! 1 V01 *-01 1000 3000 Particle S i ze ( ^m) -Ant - Whole - Average, Monday, October 06, 2003 11 01 08 A M Size(ijm) rVolumeln% 0010 000 0011 •; •'• .0.00 0013 0.00 •' -0.014 0.00 ' 0.016 .. 0.00 0.018 ' .000 0.020 0.00 ; 0.022 0.00 0.025 0.00 0.028 , 0.00 0032 0.00 . 0.035 000 0.040 , . 0.045 000 ; 000 O050 , 0 0 0 0.056 0.00 0.053 0.00 0.071 .Size (um) Volume In % 0071 .0.00 • :O079 000 , , 0089 000 • .0.100 000 .'• 0.112: 000 .... 0.126 • 000 • 0.141 0.00 .' .0.158 0.00 '. 0.178 . 0.00 -. 0.200 aoo 0.224 - -. Q00 0.251 aoo 0.282 v ' .aoo 0.316 .0.00 : ; ' 0.355 0.00 0.398 " \ '0.00 0447: . >0.00 . 0501 0.501 0.562 0.631. ,0.708 0794 0.891' 1.000 '1.122 1:259 1.413 1.565 1.778 .1.995 .2239 2512 2.818 '3.162 •3.548 aoo 000 0.00 0.00 .0.03 aoo .0.00 0.00 0.00 0.00 000 0.00 0.00 0.00 0.07 0.14 0.26 Stze (um) iVoiume In % 3548 038 3981 051 4487 065 5012 ' 079 5623 094 6310 .-1.09 : 7.079 -' 1.T) 7.943 142 / 8.913 • 10.000 178 -11:220 12589 219 . 14.125 '• ' - 2.42, .: 15849 267: 17.783 12.94 1- ,19*953 i ' ; , 3 2 4 , , 22.387 •. 356 , 25.119 Ssze(um) V^olume ln% sStze (Mm) 25119 390 427 177 828 28184 199.526 31623 223872 35481 495 524 5.45 5.55 .-• ' 552 ., ..5.35 5.02 456 251 189 39811 281 838 44668 ' ' 50.119 56.234 63096 '70.795 316.228 " 354.813 398.107 446684 501 187 .79.433 . 89.125 .. 562.341 630957 100.000 333 -, ,265 201, , 142 707946 11Z202 " 794.328 125893 891251 141.254 ' 1000.000 .158489 . 177.828 094 057 032 . 0.17: 008 005 004 .004 0.04 :0.03 -0.00 003 000 0.00 .003 : 0,00 132 A S T E R S I Z E R v _ Result Analysis Report Sample Name: SOP Name: Measured: Ant - Fine Grind - Average Monday, October 06, 2003 10:53:44 AM Sample Source & type: Measured by: Analysed: Factory contract Monday, October 06, 2003 10:53:46 AM Sample bulk lot ref: Result Source: Averaged Particle Name: Accessory Name: Analysis model: Sensitivity: Chalcopyrite Hydro 2000S (A) Generajpurpose Normal Particle Rl: Absorption: Size range: Obscuration: 1.450 0 0.020 to 2000.000 um •16:22 % Dispersant Name: Dispersant Rl: Weighted Residual: Result Emulation: Water 1.330 1.495 % Off Concentration: Span: Uniformity: Result units: 0.0279 %Vol 2.059 0:642 " Volume Specific Surface Area: Surface Weighted Mean D[3,2]: Vol. Weighted Mean D[4,3]: 0.476 m'lg 12.597 , um 20:816 um d(0.1): 6.136 d(0.5): 16.861 d(0.9): 40.855 8 7 6 5 i A 3 r 2 | 1 I Particle Size Distribution L J J LU. > i 01 0.1 10 Particle S i ze (Mm) 100 1000 3000 -Ant - Fine Grind - Average, Monday, October 06, 2003 10.53 44 A M Size(um} Volume In % 0010 • "'0.00 O011-000 0013 0.00 . .0.014 i* 0.00 0016 '." 0.00 0.018 0.00 0020 000 O022 : o.oo Q025 0.00 O028 ooo ' .0032 0.00 0.035 0.00 0040 j-0.00 0.045 . 000 -0.050 000 . 0.056 :o.oo 0063 000 O071, 0071 ,0.079 0.089 0.100 . 0.112 <0,126 0141 0.158 0.178 •0.200 .0.224 . 0.251 .0.282 0.316 .0.355 0.398 :0447: 0.501 O.OO 'ooo ooo 000 0.00 ".' aoo • 000 • 0.00 .. aoo • o.oo o.oo • o.oo aoo . 0.00 • ,0.00 ooo ooo Size (Mm) Volume In % 0501 • .0.00 '0552: 000 0.631: '.-•"'. ooo - " 0.708" ,  ,0.00 0.794' 000 - • • aoo 1:000 , aoo '-. '.1.122 ;. - • aoo 1:259 '""--. 0.00 1.413 * aoo •' 1.585 . . 0.00 1:778 '. 0.00 : 1.995^  aoi 2239 ' • ' 0.10 • 2512 - • 2818 . 0.2B 0.52 3162 -.' 0.77 3.548 Size (Mm): 3548 1981 4467 5012 5623 '. ' 6:310 .;7.079 7943 8913 •10000 11.220 v 12589 .' 14.125 15849 "17.783 .19.953 22387 . 25.119 ,1.09 142 1r78 216 256 298 342 .388 '4.34 •4.80 5.22 •6.59 5.88 -6.06 -6.12 6,03 , 580 Size(pm) 25119 28184 31623 35481 39811 "44668 60119 56.234 •63.096 70795 79.433 '89.125" 100.000 112202 -"'125.893 -.141.254 544 4.95 437 ' 373 307 244 185 135 0.92 059 .038 014 001 aoo ooo aoo •' ooo Size (Mm) 177828 199526 223872 251189 281838 316228 ,354.813 398107 '446.684 •501.187, .662341 630957 707.946 794.328 891.251; '1000.000 000 000 000 000 000 000 000 000 000 ' 000 000 000 000 •aoo 000 133 A S T E R S I Z E R Result Analysis Report Sample Name: SOP Name: Measured: Ros - Fine Grind - Average Monday, October 06. 2003 10:40:43 AM Sample Source & type: Measured by: Analysed: Factory contract Monday, October 06,2003 10:40:44 AM Sample bulk lot ref: Result Source: Averaged Particle Name: Accessory Name: Analysis model: Sensitivity: Chalcopyrite Hydro 2OO0S (A) General purpose Normal Particle Rl: Absorption: Size range: Obscuration: 1.450 0 0,020 to 2000.000 um 17.01 % Dispersant Name: Dispersant Rl: Weighted Residual: Result Emulation: Water 1.330 1.394 " % Off Concentration: Span : Uniformity: Result units: 0.0283 %Vol 1.925 0:6 Volume Specific Surface Area: Surface Weighted Mean D[3,2]: Vol. Weighted Mean D[4,3]: 0.492 m'/g 12.206 um 19.465. um d(0.1): 6.041 um d(0.S): 16.243 um d(0.9): 37.306 um Particle Size Distribution 1 01 0.1 10 Particle S i ze (pm) 1000 3000 -Ros - Fine Grind - Average, Monday, October.06, 2003 ; 10:40:43 A M Sizefum) Volume In* ;0.010 000 000 ••ao\3 000 • • 0.014 000 0016 000 0018 000 : 0.020 QOO 0022 ooo '0.025 000 :. ,0.023 0.00 ; 0.032 I. '.OOO : -0.O35 000 0.040 000 0.046 000 0.050 ooo 0056 • - 0.00 0063 0.00 0.071 Volume In %' 00711 . 000 ' O079 000^  0069 000 0100 ,000 ' 0112 0126 000 000 0141 600 0158 00b 0178 000 0200 000 0224 000 025V poo 0282 boo 0316 000 0356 000 0.396 -•000 . 0.447 0.00 ' ."0.501 Size(um) "0.501 - 0.562 0631 0706 0794 t' 0891 '1OOO 1 122 1.259 . M1413 ! -1.585 M778 . .1.995 . 2.239 2.512 '2.818 3162 . 3548 ' 0.00 •0.00 " 000 ",00a "000 000 ooo 000 ,000 000 'obo -' :0.00 001 015 o x ;0.58 -084 Size(um) Volume ln,% »,"3548 s 113 3981 . 4467 145 179 5012 216 5623 2.56 6310 7079 •2.99 347 7943 3.97 8913 10000 . 449 502 11220 550 . :12.589 593 ' 14.125 625 - , 15.843 - 644 . '17.783 .19.953 648 . 631 '.' 22.387' 25.119 'Stze(um) Volume In % 25119 550 28184 31623 419 35.481 245 39811 44668 _27f ' 204 50119 145 58234 63096 096 059 70 795 035 79433 005 89125 000 100.000 000 112302 125893 : 000 . 000 141354 »v .0.00 15^489 0.00 ..'177.828 Size (urn) Volume In % .177.828 000 199.526 " ' 0 0 0 '.,223.872 , 000 SV 251'.189 OOO 281638 - 000 316228 ,000 354813 ; , 0 0 0 398107 «•> 000 ' 446.684 - "-000 501:167 000 !• 582.341-••' '_' . 000 630.957 000 707.946 .,784.328 a'oo - .'000 891.251 000 •1003.000 134 Appendix 4 XRD Spectra Antamina XRD Profile 10 20 30 40 SO 60 7C 2-Theta - Scale Rosario XRD Profile 3400 - . 3300 ~ 3200 -4 10 20 30 40 50 60 2-Theta - Scale 135 Appendix 5 Methods A. Free Acid Titration B . Cerium Titration C. Light's Solution Preparation 136 Appendix 5A. Free Acid Titration (with oxalate masking) The purpose of this titration was to determine the free acid concentration of the feed liquor. Potassium oxalate (K2C2O4.H2O) was used to complex metal ions that would normally contribute protons (H + ) , via complexation with hydroxyl ions (OH") in water, e.g. C u 2 + + 4(OH") -> Cu(OH) 4 2 " . Method used 1. In a beaker containing ~65mL DI water, add a 1 or 2mL aliquot of the feed liquor. 2. A d d 25mL Potassium Oxalate solution (280g K o x . H 2 0 in 1L) 3. Measure the p H of the solution before and after addition of K(ox) . 4. Titrate solution with 0 .1M N a O H (accurately determined) and record p H changes with volume added. 5. Plot p H vs volume N a O H added. The mid-point of the steepest part of the curve indicates the volume required to neutralize the free acid. This volume can be used to determine moles of free acid in the aliquot used, as shown in the following reaction. OH~ + H + - * H 2 0 (A5.1) The results of titrations performed at U B C , shown in the following table, indicate that for all but H3 and H4 the results were in general agreement. There is still some slight variation from the IPL data but as the concentration of the feed liquor is extremely high in copper, iron and calcium chloride this may impact on the method and output reproducibility. Example titration curves are provided for experiment H7. 137 Table 34 Comparing IPL free acid results with test titration value Expt IPL result UBC titrations -(g/LHCl) averages given (g/L HC1) H I 5.74 Not performed H2 5.28 Not performed H3 12.97 4.06 H4 12.53 4.06 H5 2.71 3.04 H6 1.94 2.19 H7 3.01 2.96 H8 3.69 3.14 H9 2.35 2.55 G l 0.72 Not performed G2 10.02 Not performed G3 <0.01 <0.1 Example free acid titration curves for experiment H7 ( l m L and 2mL aliquots) Aliquot = 1 m L Vol (mL) pH 7.6 6.73 7.8 6.85 8.0 6.97 8.2 7.21 8.5 7.67 8.7 7.92 9.0 8.13 Calculation: Midpoint of steepest section (between points in bold): V o l = 8.35mL Titration V o l = 8.35-7.6 = 0.75mL M o l N a O H = M o l HC1 = 0 .1M x 0.75mL = 0.075mmol [HC1] = 0.075mmol/0.001L = 0.075M = 2.73g/L 138 lmL aliquot pH 8.5 8 7.5 7 6.5 6 Midpoint (8.35,7.4) 7.0 7.5 8.0 8.5 9.0 Vol NaOH O.lM(mL) 9.5 Aliquot = 2mL Vol (mL) pH 9.5 6.21 9.7 6.26 9.9 6.35 10.2 6.46 10.5 6.58 10.7 6.75 11.1 7.05 11.4 7.44 11.7 7.69 11.9 7.80 12.2 7.94 12.5 8.00 12.7 8.05 13.0 8.12 13.6 8.21 14.2 8.27 15.3 8.36 Calculation: Midpoint of steepest section (between points in bold): Vol=11.25mL Titration Vol = 11.25-9.5 = 1.75mL Mol NaOH = Mol HC1 = 0.1M x 1.75mL = 0.175mmol [HC1] =.0.175mmoV0.002L = 0.175M = 3.19g/L 2mL aliquot 9.00 8.50 8.00 7.50 pH 7.00 6.50 6.00 5.50 5.00 Midpoint (11.25,7.25) 7.0 9.0 11.0 13.0 15.0 Vol NaOH 0.1M (mL) 17.0 Average of two free acid titrations = (2.73 + 3.19)/2 = 2.96 g/L HC1 139 Appendix 5B Cerium Titration (for reduced species determination) This titration was used to determine reduced species in the goethite feed liquor and in the final L S I (PLS) and L S 2 exit liquors of selected experiments. Certified acidic cerium (Ce 4 + ) sulfate solution (0.1M) was used as the titrant, and as the C e 4 + / C e 3 + redox potential is significantly high (at +1.61V) it has the ability to oxidize many reduced metals ions, e.g. C u + or Fe 2 + , via the following reactions. C e 4 + + C u + C e 3 + + C u 2 + (A5.2) C e 4 + + F e 2 + ^ C e 3 + + F e 3 + (A5.3) 2 C e 4 + + F e 2 + + C u + -> 2 C e 3 + + F e 3 + + C u 2 * (A5.4) Method used 1. A 0.5, 1 or 2mL aliquot was added to a beaker containing 80-100mL DI water. 2. A few drops of H C I were then added, i f required, and a few drops of Ferroin indicator. 3. The change in redox potential was then recorded as 0 .1M C e 4 + titrant was introduced into the solution. 4. The endpoint was detected by a change in color of the indicator as well as a sharp rise in the redox potential, i.e. >800mVA g/Agci-5. The concentration of total reduced species was determined by calculating the moles of C e 4 + reacted using the titration volume, and then dividing by the aliquot of test solution added. 140 Example cerium titration curve for experiment H7 - Final PLS liquor Aliquot = 0.5mL Calculation: Titration Vol = 15.0-12.0mL = 13.0mL Mol NaOH = Mol HCI = 0.1M x 13.0mL = 1.30mmol [HCI] = 1.30mmol/0.0005L = 2.60M Total Reduced Species V o l (mL) E (mV A g / A gci) 2.0 162 2.5 176 2.7 181 3.2 190 3.5 199 3.9 209 4.3 227 4.5 248 4.8 320 5.4 389 5.6 405 6.2 415 6.5 421 6.8 426 7.1 430 7.4 434 7.7 438 8.2 443 9.0 450 9.3 453 10.1 460 11.1 469 12.3 481 12.6 486 13.1 493 13.6 502 14.0 512 14.4 523 14.9 553 15.0 endpoint (>800mV) Cerium titration H7 - Final PLS liquor - 0.5mL aliquot U en < > > 600 550 500 450 400 350 300 250 200 150 100 0.0 2.0 4.0 6.0 8.0 10.0 12.0 Vol 0.1 M Cerium (mL) 14.0 16.0 141 Appendix 5C.Light's Solution Preparation Redox probes cannot be recalibrated and thus performance must be monitored throughout the duration of its lifespan. A solution of known potential, such as Light 's solution, can be used to check probe performance. Dissolve 39.2lg of ferrous ammonium sulfate-6-hydrate and 48.22g of ferric ammonium sulfate-12-hydrate in 800mL of DI water. Then add 56.2ml of concentrated sulfuric acid and adjust volume to 1 liter with DI water. This solution has a p H near 0. It is a corrosive strong acid solution and should be handled with care and kept stored away from light. For a stable reading, the redox electrode should be placed into stirred light's solution for at least 30 minutes for equilibrium to be established. The value that should be obtained for A g / A g C l (sat.KCl) probes is +476mV and Calomel (sat.KCl) probes is +430mV at 25C. 142 Appendix 6 Experimental Logs A . Measurements (Temperature, p H and E) B . Samples C. Mass and Volume Transfers 143 Date: Thursday 25 September 2003 Experiment#18 (H9) Continuous 18h Countercurrent Leach, Fine Grind, No Air to LS2,135g/L, 95C, RT 3h. ROSARIO Concentrate Light's Soln: Start= Concentrate fine grind filtered, dried and rolled with SS rolling pin. I End= Leach Solution: 50g/L Cu2+, 80g/L Fe3+, 110g/L CaCI2, -3g/L HCI | Average= RT - C; E(mV/AgAgCI)= pH = Time Leach time LS1 LS2 Comments T(C) h (mV Ag/AgCI) PH T(C) t (mVAg/AgCI) pH I. 0 1.5h 2.75h A1.10toLS1 15mL 3h Bench pH -Transfer time = II. 0 1.5h 2.75h A110toLS1 15mL 3h Bench pH -Transfer time = III. 0 1.5h 2.75h A110toLS1 15mL 3h Bench pH -Transfer time = IV. 0 1.5h 2.75h A110toLS1 15mL 3h Bench pH -Transfer time = V. 0 1.5h 2.75h A110toLS1 15mL 3h Bench pH -Transfer time = VI. 0 1.5h 3h Bench pH -144 c |p's, let o O T3 <N CM CM C N T T o [M M i - CM C N C N C N C N C N O ^ < C N o oo CM X Q- 0_ co m oo o. o D . o CL o CNI X o C N X o CM X o CM X o CM X o CM X a) E E o O E o o CM E CM O X E CM o X _l E LO E E CM O X _l E i n E o o CM E CM O X _l E IT) O X E o o CM E CM O X 1° o o X E o o CN E CM 9 o o E CM O X o X 11 9 E o E o oo II a> I E o C O E o C O E o C O 0) CL E m C O I CM m oo CL E JD CO m C O m C O LO m 00 a) CL E 03 c Z SI 3 a o 3 CL _C •a c ra o t' ra (/) fl 00 C N I 00 CM 00 CO CM CM CM 00 < c o l 00 C O 52 CO < C O 4H 5 C O II LO < 00 CM I 00 152. C O >-O Q C O II Q 5 II 00 _l 0. <5 c 0) E t_ a> a x LU ra a _i t ra u ra 0) —j t ra t ra (0 ra .2 in c n C D 3 Feed Liquor 3 1L CuFeS2= 135fl . , . _ L _ L <0mL Sample Volume for L S 1 ° \ CuFeS2° 135g Volume for LS1= Transfer Vnliimn fnr I S1a Transfer Vnlumft fnr LSI: A11015mL LS1 \ Volume for LS1 l80njL Sample Experiment: Date : Page: Feed Liquor 3 1L C u F e S 2 3 135g LS2 ,  i  ° I ( 80mL Sample Pressure Filter LR wet cake=> vol. llquor?= washed and d r i e d 3 Pressure Filter \.R wet caken vol. liquor?= washed and dried= 80mL Sample Pressure Filter rLR wet cake= vol. llquor?= washed and drled= BOmL Sample Pressure Filter i . LR wet cake= vol. l iquor? 3 washed and dried* ,80ml. Sargple Pressure Filter Feed Liquor 3 1L T LR wet c a k e 3 vol. liquor?= washed and d r i e d 3 80mL ^ample Pressure Filter Pressure Filter • LR wet c a k e 3 vol. l iquor? 3 washed and dr ied 3 LR wet cake= vol. l iquor? 3 washed and dr ied 3 146 Appendix 7 Calculations A . Copper and iron extraction B . Required oxidant concentration for hematite feed liquor C. Oxygen addition to LS2 D . Theoretical reduced species 147 Appendix 7A. Copper and iron extraction The extraction of copper and iron from the concentrate was calculated in the following manner. A n example for copper is provided for experiment H5 (Antamina p90=41um, 3h, 95C). C u % in feed = 28.8%, C u % in L S residue = 1.96% Amount of copper in feed = 135g x 28.8% = 38.88g Final mass of LS2 residue (80mL slurry sample and bulk pressure filter sample) = 3.1 +37.8 = 40.9g Copper in final L S 2 residue = 40.9g x 1.96% = 0.80g Extraction of Copper = (copper in feed - copper in L S 2 residue)/ (copper in feed) x 100% = (38.88 - 0.80)/38.88 x 100 = 97.9% This calculation does not take into account the small residue sample taken from L S 1, which may decrease the lower extraction values by no more than 1-2%. Appendix 7B. Required oxidant concentration for hematite feed liquor Calculation for determining i f 70g/L ferric (Fe 3 + ) and 40g/L cupric (Cu 2 + ) in the feed liquor is sufficient for leaching 100% of CuFeS2 in the Antamina concentrate feed: 28.8% C u in 135g feed concentrate = 38.9g C u = 0.61mol If all o f the copper is contained in CuFeS2 this implies 0.61mol CuFeS2. The ferric and cupric leach reactions are as follows: CuFeS 2 + 4 F e C l 3 -» 5FeCl 2 + C u C l 2 + 2S (A7.1) CuFeS 2 + 3 C u C l 2 -i> 4 C u C l + F e C l 2 + 2S (A7.2) If Fe 3 + = 70g/L = 1.25 mol in I L , then ferric can leach 1.25/4 = 0.31mol CuFeS 2 and produce 0.31 mol C u 2 + . If C u 2 + is 40g/L (0.63 mol in IL) and is combined with the 0.31mol C u 2 + produced by the ferric leach, this amounts to C u 2 + = 0.94mol in I L . This total C u 2 + amount, 0.94mol, has the ability to leach 0.94/3 = 0.31mol CuFeS 2 . The total amount of CuFeS 2 that can be leached by 70g/L F e 3 + and 40g/L C u 2 + is therefore 0.31 + 0.31 = 0.62mol CuFeS 2 , which is 0.01 mol greater than the amount of CuFeS 2 delivered in the Antamina concentrate. 148 Appendix 7 C . Calculation for oxygen addition to L S 2 (Goethite model) Delivery of oxygen to LS2 is required in the goethite model experiments to regenerate cupric and ferric species. The following calculation assisted in determining an appropriate flow rate for air delivery. . Leach time = 2h = 120min A i r = assume 21% oxygen Oxygen requirement (determined from the flowsheet mass balance) per 135 g pure chalcopyrite feed = 17.65g = 0.55mol Moles of air required = 0.55mol x 100/21 = 2.62 mol If 1 mol gas approximates 24.5L at room temperature (298K), 2.62 mol air = 64.2L Over two hours the flow rate to deliver a stoichiometric quantity of oxygen = 64.2L x lOOOmL/L / 120min = 535 mL/min To ensure a sufficient amount of oxygen is delivered to L S 2 a 5-10 times excess should be considered, to overcome inefficiencies caused by transfer and solubility limitations. This would mean an airflow rate of 2.7L/min or 5.3L/min. For experiment G2, a flow rate of 2.5L/min was used, a 4.7 times excess. Although it would be better to have a 10 times excess this would have lowered the temperature of L S 2 too much, as even this rate depressed the temperature by 4C. For experiment G3 , almost pure oxygen (98%) gas was delivered at a flow rate of 2.5L/min. A s this gas is approximately 98/21 times more concentrated than the air and the residence time was extended from 2 to 3 hours, the actual excess of oxygen delivered is close to x33. 149 Appendix 7D. Theoretical reduced species An example calculation follows for determining the theoretical reduced species concentration for the exit PLS in experiment H2 (75um, 2h, 95C), assuming a 1L volume of feed liquor and PLS: Feed CuFeS2 = 0.612mol If copper extraction = 87%, then (assuming this is equivalent to CuFeS2 extraction) the amount of reacted CuFeS2 = 0.87 x 0.612 = 0.532mol The total reduced species produced via CuFeS2 leaching (i.e. 5mol per molCuFeS2) = 5 x 0.532mol = 2.66mol Based on 1L of feed liquor (Cu2+= 0.743 M; Fe3+= 1.431 M) the theoretical total species would be = Cu and Fe from CuFeS2 (0.532 x 2) + Cu and Fe in the feed (0.743 + 1.431) = 3.24M 150 Appendix 8 Mass balances for experimental accountability 151 Experiment H1 Sample guide: LS1 = M, LS2 = N Cu and Fe Input Solids Description wgt (g) Cu (%) Fe (%) Cu(g) ... Fe (g) Concentrate 6x 135g 810.0 28.8 28.4 233.28 230.04 Solution Vol (L) [Cu] (g/L) 47.8 [Fe](g/L). Cu (g) Fe (g) Feed liquor 6 x 1 L 6.0 76.72 286.56 460.32 Total Input | 519.84 690.36 I Input per 135g CuFeS2 Cu and Fe Output Cu = 38.88 g Samples - balance solution reading with residue - record amounts (vol or wgt) and concentration Fe= 38.34 9 Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe(g) . Cu % F e % 7N1S bulk and sample 72.2 21 22.4767 15.16 16.23 61.0 57.7 7N2S bulk and sample 46 19 24 8.74 11.04 77.5 71.2 7N3S bulk and sample 51.3 16 20 8.21 10.26 78.9 73.2 7N4S bulk and sample 73.8 16.8 18.4174 12.40 13.59 68.1 64.5 7N5S bulk and sample 44.5 17.47 22.12 7.77 9.85 80.0 74.3 7M5S bulk and sample 87.6 18 23 15.77 20.15 59.4 47.4 Solutions Vol (L) [Cu] (g/L) [Fe](g/L) Cu (g) Fe (g) 7M1 bulk and sample 0.84 65 95.1 54.60 79.88 7M2 bulk and sample 0.74 86.04 113.77 63.67 84.19 7M3 bulk and sample 0.82 71.83 98.14 58.90 80.47 7M4 bulk and sample 0.88 76.9 107.6 67.67 94.69 7M5 bulk and sample 1.19 74.09 101.69 88.17 121.01 7N5 bulk and sample 1.13 54.06 85.19 61.09 96.26 Wash 1.1 3.45 4.8 3.80 5.28 A110 4x10mL 0.04 Estimated loss from samples not analyzed (see box below) 23.58 34.84 Total Vol PLS + samples + flocc Total Output 489.52 677.74 = 5.64 I Total Expected Cu and Fe 519.84 690.36 in concentrate and solution % Recovery 94.171 98.17 ** Note this includes the estimated copper and iron in samples not analyzed (see box below) Unaccounted for mass output: 7M1-M4 residues = 33.2g 7N1-4 solutions = 320mL I Estimate additional mass of this output here: Estimate average LS1 residue Cu and Fe % (from7M5) Thus Est % Total g in 33.2g Cu 18 5.976 Fe 23 7.636 Estimate average LS2 liquor Cu and Fe g/L (from7N5) Est g/L Total g in 320mL Cu 55 17.6 Fe 85 27.2 Total extra output (g) Cu 23.576 Fe 34.836 152 Experiment H2 Sample guide: LS1 = K, LS2 = L. Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 6 x135g 810.0 28.8 28.4 233.28 230.04 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe(g) Feed liquor 6 x 1 L 6.0 47.2 79.5 283.48 476.84 Total Input 516.761 706.88 Input per135g CuFeS2 Cu and Fe Output Cu = 38.88 g Fe= 38.34 9 Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) C u % F e % 9 L 1 S bulk and sample 56.98 17.00 19.61 9.69 11.18 75.1 70.9 9 L 2 S bulk and sample 34.8 13.41 15.81 4.67 5.50 88.0 85.7 9 L 3 S bulk and sample 53.13 11.52 14.98 6.12 7.96 84.3 79.2 9 L 4 S bulk and sample 40.9 7.62 11.10 3.12 4.54 92.0 88.2 9 L 5 S bulk and sample 62.6 8.13 12.44 5.09 7.79 86.9 79.7 9 K 1 S sample 5 16.16 19.72 0.81 0.99 9 K 2 S sample 7.1 23.17 24.55 1.64 1.74 9 K 3 S sample 5.2 16.28 19.11 0.85 0.99 9 K 4 S sample 5.4 18.76 20.44 1.01 1.10 9 K 5 S bulk and sample 79.3 17.85 19.57 14.16 15.52 63.6 59.5 Solutions Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) 9K1 bulk and sample 0.88 74.05 108.59 65.16 95.56 9K2 bulk and sample 0.66 86.16 120.15 56.87 79.30 9 K 3 bulk and sample 0.83 78.67 111.14 65.30 92.24 9K4 bulk and sample 0.97 81.76 114.75 79.31 111.30 9 K 5 bulk and sample 1.15 81.75 114.28 94.01 131.42 9L1 sample 0.08 75.64 110.40 6.05 8.83 9L2 sample 0.08 54.25 89.27 4.34 7.14 9L3 sample 0.08 66.32 100.84 5.31 8.07 9L4 sample 0.08 58.47 92.08 4.68 7.37 9L5 bulk and sample 1.21 60.11 93.09 72.73 112.64 Wash 3.1 4.17 6.35 12.91 19.67 A110 4x1 OmL 0.04 Sum of volumes = 6.06 Total Output I 513.82 730.85 I I I . .1 Total Expected Cu and Fe I 516.76 706.88 in concentrate and solution | I | % Recoveryl 99.43 103.39 153 Experiment H3 Sample guide: LS1 = C, LS2 = D. Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 7 x 1 3 5 g 945.0 28.8 28.4 272.16 268.38 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 7 x 1 L ' 7.0 52.9 87.2 370.321 610.106 Total Input 642.48 878.49 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.88 g Fe= 38.34 9 Extraction Solids Description wgt(g) Cu (%) Fe (%) Cu (g) Fe (g) Cu % F e % 11D1S bulk and sample 55.8 20.61 22.68 11.50 12.65 70.4 67.0 11D2S bulk and sample 49 17.66 20.36 8.65 9.98 77.7 74.0 11D3S bulk and sample 50.5 14.29 17.98 7.22 9.08 81.4 76.3 11D4S bulk and sample 44.8 15.22 19.31 6.82 8.65 82.5 77.4 11D5S bulk and sample 49.9 12.76 17.00 6.37 8.48 83.6 77.9 11D6S bulk and sample 57.7 11.32 16.24 6.53 9.37 83.2 75.6 1 1 C 1 S sample 4.2 19.44 22.30 0.82 0.94 1 1 C 2 S sample 5.9 25.30 25.86 1.49 1.53 1 1 C 3 S sample 4.9 19.93 22.34 0.98 1.09 1 1 C 4 S sample 4.9 22.41 24.06 1.10 1.18 1 1 C 5 S sample 4.9 19.03 21.22 0.93 1.04 1 1 C 6 S bulk and sample 88.9 20.73 22.02 18.43 19.58 52.6 48.9 Solutions Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) 11C1 bulk and sample 0.83 74.03 110.70 61.44 91.88 11C2 bulk and sample 0.75 90.71 130.22 68.03 97.67 11C3 bulk and sample 0.91 78.88 116.05 71.78 105.60 11C4 bulk and sample 0.87 85.34 123.39 74.25 107.35 1 1 C 5 bulk and sample 0.85 54.54 123.83 46.36 105.26 11C6 bulk and sample 1.09 86.28 125.09 94.04 136.34 11D1 sample 0.08 76.16 112.14 6.09 8.97 11D2 sample 0.08 57.72 93.34 4.62 7.47 11D3 sample 0.08 66.56 105.49 5.32 8.44 11D4 sample 0.08 61.37 96.52 4.91 7.72 11D5 sample 0.08 64.75 102.70 5.18 8.22 11D6 bulk and sample 1.08 63.03 100.70 68.07 108.75 Wash 3.75 5.12 7.68 19.19 28.81 A 1 1 0 5x10mL 0.05 Sum of volumes = 6.83 Total Output 600.13 906.05 Gust PLS+sample not inside LS1) I I I Total Expected Cu and Fe 642.48 878.49 in concentrate and solution I % Recoveryl 93.41 103.14 154 Experiment H4 Sample guide: LS1 = M, LS2 = N. Cu and Fe Input ,. Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 7 x 1 3 5 g 945.0 28.8 28.4 272.16 268.38 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 7 x 1 L 7.0 55.8 90.5 390.733 633.325 Total Input , 662.89 . 901.71 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.88 g Fe= 38.34 g Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) C u % F e % 12N1S bulk and sample 42.7 13.32 18.37 5.69 7.85 85.4 79.5 12N2S bulk and sample 35.9 7.10 12.84 2.55 4.61 93.4 88.0 12N3S bulk and sample 48.5 8.71 14.97 4.22 7.26 89.1 81.1 12N4S bulk and sample 40.6 5.78 12.39 2.35 5.03 94.0 86.9 1 2 N 5 S bulk and sample 46.9 7.31 13.80 3.43 6.47 91.2 83.1 1 2 N 6 S bulk and sample 45 6.14 12.48 2.76 5.61 92.9 85.4 1 2 M 1 S sample 3.4 13.34 17.38 0.45 0.59 1 2 M 2 S sample 9.7 26.99 29.65 2.62 2.88 1 2 M 3 S sample 4.2 16.67 19.61 0.70 0.82 1 2 M 4 S sample 4.2 20.56 22.23 0.86 0.93 1 2 M 5 S sample 4 17.59 20.06 0.70 0.80 1 2 M 6 S bulk and sample 46.6 21.37 24.33 9.96 11.34 74.4 70.4 Solutions Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) 12M1 bulk and sample 0.81 86.46 122.65 70.04 99.35 12M2 bulk and sample 0.66 102.52 133.30 67.66 87.98 12M3 bulk and sample 0.85 91.95 125.08 78.16 106.32 12M4 bulk and sample 0.9 93.27 125.93 83.94 113.34 12M5 bulk and sample 0.92 91.94 125.02 84.58 115.02 12M6 bulk and sample 1.1 94.59 127.92 104.05 140.71 12N1 sample 0.08 88.81 129.48 7.10 10.36 12N2 sample 0.08 66.13 103.07 5.29 8.25 12N3 sample 0.08 76.59 113.30 6.13 9.06 12N4 sample 0.08 68.18 104.11 5.45 8.33 12N5 sample 0.08 72.88 107.89 5.83 8.63 12N6 bulk and sample 1.2 68.09 102.95 81.70 123.53 Wash 3.5 5.80 8.37 20.29 29.28 A110 5x1 OmL 0.05 Sum of volumes = 6.89 Total Output 656.53 914.36 (just PLS+sample not inside LS1) Total Expected Cu and Fe 662.89 901.71 in concentrate and solution j % Recovery 99.04 101.40 155 Experiment H5 Sample guide: LS1 = G, LS2 = H. Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 7 x 1 3 5 g 945.0 28.5 28.7 269.19317 271.367 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 7 x 1 L 7.0 45.0 75.8 314.979 530.509 Total Input 584.17 801.88 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.45617 g Fe= 38.76667 9 Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) C u % F e % 13H1S bulk and sample 46.6 11.61 16.01 5.41 7.46 85.9 80.7 13H2S bulk and sample 15.2 1.40 7.73 0.21 1.17 99.4 97.0 13H3S bulk and sample 26.8 3.56 9.87 0.95 2.65 97.5 93.2 13H4S bulk and sample 36.8 0.94 7.65 0.35 2.81 99.1 92.7 13H5S bulk and sample 49 2.50 8.59 1.23 4.21 96.8 89.1 13H6S bulk and sample 40.9 1.96 8.50 0.80 3.48 97.9 91.0 1 3 G 1 S sample 4.6 11.83 15.87 0.54 0.73 1 3 G 2 S sample 11.2 27.80 25.86 3.11 2.90 1 3 G 3 S sample 6.6 18.31 20.66 1.21 1.36 1 3 G 4 S sample 10 22.71 23.14 2.27 2.31 1 3 G 5 S sample 6.5 22.23 23.44 1.45 1.52 1 3 G 6 S bulk and sample 113.9 25.03 24.73 28.50 28.17 25.9 27.3 Solutions Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) 13G1 bulk and sample 0.83 77.84 106.69 64.61 88.56 13G2 bulk and sample 0.67 80.44 114.40 53.89 76.65 13G3 bulk and sample 0.83 81.18 110.75 67.38 91.92 13G4 bulk and sample 0.83 82.79 111.76 68.71 92.76 13G5 bulk and sample 0.83 85.15 115.20 70.67 95.62 13G6 bulk and sample 1.28 86.08 118.04 110.18 151.09 13H1 sample 0.08 79.68 109.12 6.37 8.73 13H2 sample 0.08 57.19 85.11 4.58 6.81 13H3 sample 0.08 77.43 105.87 6.19 8.47 13H4 sample 0.08 59.86 87.60 4.79 7.01 13H5 sample 0.08 76.45 106.07 6.12 8.49 13H6 bulk and sample 1 71.08 100.25 71.08 100.25 Wash 6.94 10.11 0.00 0.00 A110 5x1 OmL 0.05 | Sum of volumes = 6.72 Total Output | 580.61 795.12 (just PLS+sample not inside LS1) I | I Total Expected Cu and Fe 584.17 801.88 | in concentrate and solution I | %Recovery| 99.39 99.16 156 Experiment H6 Sample guide: LS1 = V, LS2 = W. Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 7 x 1 3 5 g 945.0 28.5 28.7 269.19317 271.367 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 7 x 1 L 7.0 41.1 70.2 287.602 491.54 Total Input I | 556.80 762.91 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.45617 9 Fe= 38.76667 9 Extraction Solids Description Wgt(g) Cu (%) Fe (%) Cu (g) Fe (g) C u % F e % 1 4 W 1 S bulk and sample 47.7 11.95 15.70 5.70 7.49 85.2 80.7 1 4 W 2 S bulk and sample 34.39 1.86 8.04 0.64 2.76 98.3 92.9 1 4 W 3 S bulk and sample 56.6 5.74 10.69 3.25 6.05 91.6 84.4 1 4 W 4 S bulk and sample 43.5 3.05 9.18 1.32 3.99 96.6 89.7 1 4 W 5 S bulk and sample 52 5.76 10.90 2.99 5.67 92.2 85.4 1 4 W 6 S bulk and sample 46.6 3.69 9.59 1.72 4.47 95.5 88.5 14V1S sample 3.7 10.49 14.64 0.39 0.54 14V2S sample 10.9 27.32 25.22 2.98 2.75 14V3S sample 5.5 17.35 19.71 0.95 1.08 14V4S sample 8.6 25.99 25.05 2.24 2.15 14V5S sample 5.9 20.77 22.63 1.23 1.33 14V6S bulk and sample 100.1 25.21 24.76 25.23 24.79 34.4 36.1 Solutions Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) 14V1 bulk and sample 0.83 72.09 102.26 59.83 84.87 14V2 bulk and sample 0.62 75.13 107.06 46.58 66.38 14V3 bulk and sample 0.86 76.37 106.26 65.68 91.39 14V4 bulk and sample 0.82 73.47 98.22 60.25 80.54 14V5 bulk and sample 0.89 72.17 99.42 64.23 88.48 14V6 bulk and sample 1.13 75.27 103.88 85.05 117.38 14W1 sample 0.08 73.27 99.84 5.86 7.99 14W2 sample 0.08 52.77 79.11 4.22 6.33 14W3 sample 0.08 68.29 91.97 5.46 7.36 14W4 sample 0.08 57.37 85.08 4.59 6.81 14W5 sample 0.08 64.54 92.00 5.16 7.36 14W6 bulk and sample 1.23 60.76 87.44 74.74 107.55 Wash 3.6 5.92 8.33 21.33 29.99 A110 2x5mL+3x10mL 0.04 Sum of volumes = 6.820 Total Output 551.631 765.49 (just PLS+sample not inside LS1) I I Total Expected Cu and Fe 556.80 762.91 in concentrate and solution I % Recovery 99.07| 100.341 157 Experiment H7 sample guide: LS1 = S, LS2 = T. Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 7x 135g 945.0 28.5 28.7 269.19317 271.367 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 7x1L 7.0 47.7 80.1 333.97 560.826 Total Input I . 603.16 832.19 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.45617 g Fe= 38.76667 g Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) C u % F e % 15T1 bulk and sample 54.1 11.73 15.58 6.35 8.43 83.5 78.3 15T2 bulk and sample 37.4 2.84 8.36 1.06 3.13 97.2 91.9 15T3 bulk and sample 59.1 7.60 12.35 4.49 7.30 88.3 81.2 15T4 bulk and sample 38.8 3.80 10.52 1.48 4.08 96.2 89.5 15T5 bulk and sample 43.3 7.79 12.72 3.37 5.51 91.2 85.8 15T6 bulk and sample 43 4.03 10.20 1.73 4.39 95.5 88.7 15S1 sample 3.6 11.06 15.29 0.40 0.55 15S2 sample 9.2 26.99 25.72 2.48 2.37 15S3 sample 5.5 16.07 18.66 0.88 1.03 15S4 sample 6 23.98 24.49 1.44 1.47 15S5 sample 5.3 18.68 20.45 0.99 1.08 15S6 bulk and sample 71.5 20.65 21.77 14.76 15.57 61.6 59.8 Solutions Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) 15S1 bulk and sample 0.85 79.08 112.11 67.22 95.30 15S2 bulk and sample 0.53 84.58 117.90 44.83 62.49 15S3 bulk and sample 0.68 81.47 114.83 55.40 78.08 15S4 bulk and sample 0.89 87.66 121.39 78.01 108.04 15S5 bulk and sample 0.78 85.25 117.55 66.49 91.69 15S6 bulk and sample 1.2 81.10 112.93 97.32 135.52 15T1 sample 0.08 79.81 113.13 6.39 9.05 15T2 sample 0.08 55.91 88.41 4.47 7.07 15T3 sample 0.08 74.05 107.17 5.92 8.57 15T4 sample 0.08 61.99 94.96 4.96 7.60 15T5 sample 0.08 67.27 100.44 5.38 8.04 15T6 bulk and sample 1.2 65.42 96.92 78.51 116.30 Wash 4 5.64 7.93 22.57 31.73 A110 5x1 (toil- 0.05 Sum of volumes = 6.580 Total Output 576.91 814.36 Qust PLS+sample not inside LS1) | Total Expected Cu and Fe 603.16 832.19 in concentrate and solution I % Recovery 95.65 97.86 158 Experiment H8 Sample guide: LS1 = X, LS2 = Y. Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 7x135g 945.0 28.5 28.7 269.19317 271.367 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 7x1L 7.0 48.0 75.3 336.105 526.792 Total Input 605.30 798.16 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.45617 g Fe= 38.76667 g Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) C u % Fe% 16Y1 bulk and sample 54.8 14.68 18.53 8.05 10.16 79.1 73.8 16Y2 bulk and sample 43.3 6.66 12.02 2.88 5.21 92.5 86.6 16Y3 bulk and sample 68.3 10.12 14.49 6.91 9.90 82.0 74.5 16Y4 bulk and sample 47.3 6.94 13.35 3.28 6.31 91.5 83.7 16Y5 bulk and sample 49 9.36 14.44 4.58 7.08 88.1 81.7 16Y6 bulk and sample 49.1 7.85 13.36 3.85 6.56 90.0 83.1 16X1 sample 4.6 14.60 18.26 0.67 0.84 16X2 sample 8.4 26.36 27.31 2.21 2.29 16X3 sample 5.8 18.57 21.71 1.08 1.26 16X4 sample 7.1 23.13 24.72 1.64 1.75 16X5 sample 6.7 21.58 23.69 1.45 1.59 16X6 bulk and sample 87 23.02 24.93 20.03 21.69 47.9 44.1 Solutions Vol (L) [Cu] (g/L) [Fe](g/L) Cu (g) Fe (g) 16X1 bulk and sample 0.82 76.62 104.92 62.83 86.03 16X2 bulk and sample 0.67 81.63 109.15 54.69 73.13 16X3 bulk and sample 0.8 78.90 104.68 63.12 83.74 16X4 bulk and sample 0.81 81.20 107.60 65.77 87.16 16X5 bulk and sample 0.8 80.58 107.48 64.46 85.98 16X6 bulk and sample 1.04 80.44 106.45 83.66 110.71 16Y1 sample 0.08 76.48 104.64 6.12 8.37 16Y2 sample 0.08 56.55 83.30 4.52 6.66 16Y3 sample 0.08 68.52 93.92 5.48 7.51 16Y4 sample 0.08 61.31 88.36 4.90 7.07 16Y5 sample 0.08 66.49 92.02 5.32 7.36 16Y6 bulk and sample 1.25 63.82 91.53 79.77 114.41 Wash 3.85 5.76 8.03 22.18 30.90 A110 2x10mL+3x15mL 0.065 Sum of volumes = 6.655 Total Output 579.46 783.68 (just PLS+sample not inside LS1) j Total Expected Cu and Fe 605.30 798.16 in concentrate and solution | % Recovery 95.73 98.19 159 Experiment H9 Sample guide: LS1 = A, LS2 = B. Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 7x135g 945.0 26.8 26.8 252.86499 253.569 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 7x1L 7.0 50.6 82.0 354.0228 573.806 Total Input 606.89 827.38 Input per135g CuFeS2 Cu and Fe Output Cu = 36.12357 g Fe= 36.22415 g Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) C u % F e % 18B1 bulk and sample 70.1 0.67 28.18 0.47 19.75 98.7 45.5 18B2 bulk and sample 63.5 0.40 29.42 0.25 18.68 99.3 48.4 18B3 bulk and sample 59.1 0.57 28.83 0.34 17.04 99.1 53.0 18B4 bulk and sample 65.8 0.51 28.00 0.34 18.42 99.1 49.1 18B5 bulk and sample 74.3 1.46 28.62 1.09 21.27 97.0 41.3 18B6 bulk and sample 68.6 1.09 27.47 0.75 18.85 97.9 48.0 18A1 sample 5.6 0.67 29.05 0.04 1.63 18A2 sample 9.1 14.10 30.14 1.28 2.74 18A3 sample 5.9 0.86 28.58 0.05 1.69 18A4 sample 7.2 5.34 29.71 0.38 2.14 18A5 sample 6 2.14 28.09 0.13 1.69 18A6 bulk and sample 76.3 2.07 28.26 1.58 21.56 95.6 40.5 Solutions Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) 18A1 bulk and sample - 0.82 86.81 96.35 71.19 79.00 18A2 bulk and sample 0.68 106.80 104.54 72.62 71.09 18A3 bulk and sample 0.88 90.53 102.09 79.66 89.84 18A4 bulk and sample 0.78 99.81 104.71 77.85 81.67 18A5 bulk and sample 0.9 88.23 100.93 79.41 90.84 18A6 bulk and sample 0.98 91.68 100.48 89.84 98.47 18B1 sample 0.08 87.20 100.22 6.98 8.02 18B2 sample 0.08 55.59 86.74 4.45 6.94 18B3 sample 0.08 68.29 90.82 5.46 7.27 18B4 sample 0.08 55.46 83.54 4.44 6.68 18B5 sample 0.08 61.38 91.45 4.91 7.32 18B6 bulk and sample 1.1 55.13 84.91 60.65 93.40 Wash 4.9 2.97 3.92 14.55 19.22 A110 5x15mL 0.075 Sum of volumes = 6.615 Total Output 578.70 805.21 (just PLS+sample not inside LS1) I I Total Expected Cu and Fe 606.89 827.38 in concentrate and solution I % Recovery 95.36 97.32 160 Experiment G1 Sample guide: LS1 = P, LS2 = Q Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 6x135g 810.0 28.8 28.4 233.28 230.04 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 6x 1L 6.0 54.1 35.0 324.35379 209.821 Total Input 557.63 439.86 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.88 g Fe= 38.34 9 Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) C u % Fe% 8Q1S bulk and sample 90.8 15.89 31.72 14.43 28.80 62.9 24.9 8Q2S bulk and sample 60.8 11.38 32.28 6.92 19.62 82.2 48.8 8Q3S bulk and sample 97.1 13.99 31.27 13.58 30.37 65.1 20.8 8Q4S bulk and sample 102.8 14.32 30.32 14.72 31.17 62.1 18.7 8Q5S bulk and sample 112.4 13.87 29.80 15.59 33.49 59.9 12.6 8P1S sample 5 24.32 24.72 1.22 1.24 8P2S sample 7.3 26.69 27.14 1.95 1.98 8P3S sample 6.4 25.81 26.81 1.65 1.72 8P4S sample 5 26.42 27.33 1.32 1.37 8P5S bulk and sample 90.4 25.49 25.42 23.05 22.98 40.7 40.1 Solutions Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) 8P1 bulk and sample 0.85 62.66 41.61 53.26 35.37 8P2 bulk and sample 0.68 71.25 34.34 48.45 23.35 8P3 bulk and sample 0.78 69.05 35.45 53.86 27.65 8P4 bulk and sample 0.93 68.54 35.02 63.74 32.57 8P5 bulk and sample 1.05 51.99 25.36 54.59 26.63 8Q1 sample 0.08 71.22 31.62 5.70 2.53 8Q2 sample 0.08 60.23 27.78 4.82 2.22 8Q3 sample 0.08 69.10 32.34 5.53 2.59 8Q4 sample 0.08 68.77 31.45 5.50 2.52 8Q5 bulk and sample 1.18 67.32 29.55 79.43 34.87 Wash 3.57 6.05 4.27 21.61 15.25 SUM of volumes= 5.79 Total Output | 490.92 378.29 I I I Total Expected Cu and Fe 557.63 439.86 in concentrate and solution I | % Recoveryl 88.04 86.00 161 Experiment G2 Sample guide: LS1 = J, LS2 = K. Cu and Fe Input Solids Description Wgt(g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 6x135g 810.0 28.8 28.4 233.28 230.04 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 6x1L 6.0 55.9 34.5 335.304 207.036 Total Input 568.58 437.08 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.88 g Fe= 38.34 9 Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe(g) C u % Fe% 10K1S bulk and sample 102.2 16.25 26.60 16.61 27.18 57.3 29.1 10K2S bulk and sample 56.6 12.95 20.83 7.33 11.79 81.1 69.3 10K3S bulk and sample 35.5 10.25 21.51 3.64 7.64 90.6 80.1 10K4S bulk and sample 80.3 15.82 23.10 12.70 18.55 67.3 51.6 10K5S bulk and sample 77.8 16.55 23.17 12.87 18.03 66.9 53.0 10J1S sample 6.1 24.31 23.01 1.48 1.40 10J2S sample 9.4 30.30 27.78 2.85 2.61 10J3S sample 16.7 25.89 26.12 4.32 4.36 10J4S sample 8.4 26.90 25.30 2.26 2.12 10J5S bulk and sample 92.9 29.10 26.85 27.03 24.95 30.5 34.9 Solutions Vol (L) [Cu] (g/L) [Fe](g/L) Cu (g) Fe (g) 10J1 bulk and sample 0.81 77.40 55.10 62.69 44.63 10J2 bulk and sample 0.68 89.69 56.47 60.99 38.40 10J3 bulk and sample 0.88 84.79 56.65 74.62 49.85 10J4 bulk and sample 0.85 85.73 55.56 72.87 47.22 10J5 bulk and sample 1.08 87.95 59.73 94.98 64.51 10K1 sample 0.08 87.00 51.25 6.96 4.10 10K2 sample 0.08 71.71 44.89 5.74 3.59 10K3 sample 0.08 78.14 48.42 6.25 3.87 10K4 sample 0.08 75.67 48.43 6.05 3.87 10K5 bulk and sample 1.08 76.74 48.48 82.87 52.35 Wash 3.1 5.91 4.51 18.33 13.99 A110 4x1 OmL 0.04 Sum of volumes = 5.74 Total Output 583.45 445.02 Gust PLS+sample - not inside LS1) | Total Expected Cu and Fe 568.58 437.08 in concentrate and solution | %. Recovery 102.61 101.82| 162 Experiment G3 Sample guide: LS1 = K, LS2 = L Cu and Fe Input Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) Fe (g) Concentrate 7x135g 945.0 28.5 28.7 269.325 271.215 Solution Vol (L) [Cu] (g/L) [Fe] (g/L) Cu (g) Fe (g) Feed liquor 7x1L 7.0 47.4 30.0 332.045 209.825 Total Input 601.37 481.04 Input per 135g CuFeS2 Cu and Fe Output Cu = 38.475 g Fe= 38.745 g Extraction Solids Description Wgt (g) Cu (%) Fe (%) Cu (g) |Fe_(9) C u % Fe% 17L1 bulk and sample 134.6 1.35 43.24 1.81 58.20 . 95.3 -50.2 17L2 bulk and sample 112.9 2.88 40.79 3.25 46.06 91.6 -18.9 17L3 bulk and sample 136.9 2.59 43.52 3.55 59.58 90.8 -53.8 17L4 bulk and sample 116.3 3.72 40.04 4.33 46.57 88.8 -20.2 17L5 bulk and sample 104.5 3.93 40.12 4.10 41.93 89.3 -8.2 17L6 bulk and sample 108.2 4.04 38.42 4.37 41.57 88.6 -7.3 17K1 sample 7.1 23.40 26.11 1.66 1.85 17K2 sample 13.7 27.14 28.12 3.72 3.85 17K3 sample 8 25.18 26.84 2.01 2.15 17K4 sample 7.9 25.51 27.01 2.02 2.13 17K5 sample 8.7 24.90 26.59 2.17 2.31 17K6 bulk and sample 88.2 24.49 26.54 21.60 23.40 43.9 39.6 Solutions Vol (L) [Cu] (g/L) 60.65 [Fe] (g/L) Cu (g) Fe (g) 17K1 bulk and sample 0.825 42.46 50.04 35.03 17K2 bulk and sample 0.39 107.42 15.95 41.89 6.22 17K3 bulk and sample 0.72 91.27 25.95 65.72 18.68 17K4 bulk and sample 0.74 97.26 20.29 71.97 15.02 17K5 bulk and sample 0.755 95.46 24.44 72.07 18.45 17K6 bulk and sample 0.8 93.91 26.18 75.13 20.94 17L1 sample 0.08 98.58 0.11 7.89 0.01 17L2 sample 0.08 71.62 10.75 5.73 0.86 17L3 sample 0.08 86.55 5.52 6.92 0.44 17L4 sample 0.08 79.09 9.00 6.33 0.72 17L5 sample 0.08 76.75 9.70 6.14 0.78 17L6 bulk and sample 0.89 79.33 11.34 70.61 10.09 Wash 6 9.42 1.06 56.50 6.38 A110 5x15mL 0.075 Sum of volumes = 5.595 Total Output 591.51 463.23 (just PLS+sample not inside LS1) | I Total Expected Cu and Fe 601.37 481.04 in concentrate and solution % Recovery 98.36 96.30 163 Appendix 9 Experimental Profiles 164 Experiment H1: Coarse Antamina concentrate, 85C, 2h per leach stage Temperature vs Time 89.0 88.0 87.0 Temp (C)86.0 85.0 84.0 83.0 •LS1 -LS2 5 10 Time (h) 15 pH vs Time 5 10 Time (h) Potential, E (mV vs SCE@85C) vs Time 0 600 < 1 550 > E L U 5 Time(h) 1 0 15 Cu and Fe Concentrations vs Time 120 -100 -80 i c 1 o I 60 -c <D O 40 -o O 20 0 —A—LS1 Feg/L LS2 Fe g/L — • — L S 1 Cug/L — • — L S 2 Cug/L 0 2 4 Timl(h) 8 10 12 Cu and Fe Extraction (LS2 Residue) 77.5 tti.H 6 8 * BU.Or -Cu -Fe 15 165 Experiment H2: Coarse Antamina concentrate, 95C, 2h per leach stage 0.00 -0.10 -0.20 pH -0.30 -0.40 -0.50 -0.60 pH vs Time —•—LS1 —•—LS2 • 4 6 8 Time (h) 10 12 o < < > E LU Potential, E (mV vs Ag/AgCI@95C) vs Time 750 700 650 600 550 500 450 400 • , l ^ -1 a Lo\ - » - L S 2 4 6 8 Time (h) 10 12 Cu and Fe Concentrations vs Time Time (h) Cu and Fe Extraction (LS2 Residue) B6.9~ -•— Cu Fe 5 10 Time (h) 15 166 Experiment H3: Coarse Antamina concentrate, 85C, 3h per leach stage 800 750 400 Potential, E (mV vs Ag/AgCI@85C) vs Time 6 9 12 Time (h) 15 18 Cu and Fe Concentrations vs Time 10 Time (h) 15 20 Cu and Fe Extraction (LS2 Residue) 82,5 83.61 ^ 3 . 2 - • — C u F e 10 Time (h) 15 20 167 Experiment H4: Coarse Antamina concentrate, 95C, 3h per leach stage Temperature vs Time 100 99 98 Temp (C) 97 96 95 94 •LS1 -LS2 Bar 0 5 10 15 20 Time (h) Potential, E (mV vs Ag/AgCI@95C) vs Time 800 750 6" 700 CD ^ 650 < 600 5 10 15 Time (h) 20 Cu and Fe Concentrations vs Time 10 Time (h) Cu and Fe Extraction (LS2 Residue) 94.0—2—02.0 20 168 Experiment H5: Fine grind Antamina concentrate, 95C, 3h per leach stage Temperature vs Time 100 99 98 97 Temp 96 (C) 95 94 93 92 91 90 •LS1 -LS2 5 10 15 20 Time (h) pH vs Time Potential, E (mV vs Ag/AgCI@95C) vs Time 800 20 Cu and Fe Concentrations vs Time 5 10 Time (h) 15 20 Cu and Fe Extraction (LS2 Residue) 99.4 Q7 * 99..1 96.8 97.9 10 15 Time (h) 20 169 Experiment H6: Fine grind Antamina concentrate, 95C, 2h per leach stage pH vs Time 15 Time (h) Potential, E (mV vs Ag/AgCI@95C) vs Time -LS1 LS2 5 Time (h) 10 15 Cu and Fe Concentrations vs Time 5 T ime (h) 1 0 15 Cu and Fe Extraction L S 2 Residue 100 90 80 70 60 % 50 40 30 20 10 0 85.2 y£>^*^^t~-^+r----+ — • — C u - • - F e I I o.o 5 10 Time (h) 15 170 Experiment H7: Fine grind Antamina concentrate, 85C, 3h per leach stage 90 -89 -88 -87 -Tem 3 86. (C) 85 -84 -83 82 -81 80 -Temperature vs Time •LS1 -LS2 10 Time (h) 15 20 Potential, E (mV vs Ag/AgCI@85C) vs Time •LS1 -LS2 20 Cu and Fe Concentrations vs Time Time (h]^ 15 20 Cu and Fe % in LS2 Residue 10 Time (h) - • - C u Fe It 15 20 171 Experiment H8: Fine grind Antamina concentrate, 85C, 2h per leach stage Temperature vs Time Temp (C) 90 89 88 87 86 85 84 83 82 81 80 - • - L S I - • - L S 2 2 4 6 8 10 12 14 Time (h) Potential, E (mV vs Ag/AgCI@85C) vs Time O < cn < > > E 800 750 700 650 600 550 500 450 400 350 —•—LS1 — • — L S 2 6 8 10 12 14 Time (h) Cu and Fe Concentrations vs Time 4 x - 6 a Time (h 10 12 14 Cu and Fe Extraction in LS2 Residue -92r§-6 8 10 Time (h) 172 Experiment H9: Fine grind Rosario concentrate, 95C, 3h per leach stage Temp (C) 98 97 96 95 94 93 92 91 90 89 88 Temperature vs Time A A A A a A w • T * X L II • \ —•—LS1 —H—LS2 10 Time (h) 15 20 800 750 0 700 £ 650 £ 600 5 550 | 500 uT 450 400 350 Potential, E (mV vs Ag/AgCI@95C) vs Time —•—LS1 —•—LS2 0 2 4 6 8 10 12 14 Time (h) Cu and Fe Concentrations vs Time 10 Time (h) 15 20 173 Experiment G1: Air to LS2 (200ml/min), Coarse Antamina concentrate, 85C, 2h per leach stage Temperature vs Time 89 T 88 -87 -Temp 86 (C) <> 85 84 -83 -82 -PH pH vs Time —•—LS1 — • — L S 2 .11 ^ J w—• ™— = — II 4 6 8 10 12 Time (h) Potential, E (mV vsAg/AgCI@85C) vs Time 400 -•—LSI - • - L S 2 0 2 4 6 8 10 12 Time (h) Cu and Fe Concentrations vs Time 80 70 _ 60 3 50 I 40 2 | 30 § 20 o 10 0 • A _ _ —•—LS1 Cu —•—LS2 Cu ""^  ^^^^xjE —*—LS1 Fe — • — LS2 Fe 0 2 4 6 / u 8 10 12 Time (h) Cu and Fe % in LS2 Residue 174 Experiment G2: Air to LS2 (2.5L/min), Coarse Antamina concentrate, 95C, 2h per leach stage Temperature vs Time pH vs Time - • - L S I LS2 8 10 1B Time (h) Potential, E (mV vs Ag/AgCI@95C) vs Time 580 560 s=> 540 o S> 520 |j> 500 <n 480 > 460 ~ 440 LU 420 400 • LS1 • LS2 2 4 6 8 10 12 Time (h) Cu and Fe Concentrations vs Time 100 90 ~ 80 3 70 .2 6 0 2 50 1 40 § 30 0 20 10 0 r^y^—•— • — /s —•—• IF i A A 11 A * Mr m y if —•—LS1 Cu —•—LS2 Cu —4—LS1 Fe —•—LS2 Fe 4 6 8 Time (h) 10 12 Cu and Fe Extraction (LS2 Residue) 67.3 6 6 , 9 4 6 8 10 12 Time (h) 175 Experiment G3: 0 2 to LS2 (2.5L/min), Fine grind Antamina concentrate, 95C, 3h per leach stage pH vs Time - •—LS1 LS2 IcT* 1% • 20 Time (h) Potential, E (mV vs Ag/AgCI@95C) vs Time 350 10 Time (h) 15 20 120 Cu and Fe Concentrations vs Time 10 15 Time (h) — • — LS1 Cu LS2 Cu m * LS1 Fe m LS2 Fe 20 Cu and Fe % in LS2 Residue - • — C u • • • » 15 20 Cu and Fe Extraction (LS2 Residue) 100 95.3 91.6 90.8 88.8 89.3 88.6 -Cu Fe Time (h) 176 

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