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Analysis of the first-stage leach process at Inco Ltd.’s copper refinery Van Lier, Roy J. M. 1996-12-31

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ANALYSIS OF T H E FIRST-STAGE L E A C H PROCESS A T INCO LTD.'S COPPER REFINERY  by ROY J.M. V A N LIER M . S c , Delft University of Technology, 1993  A THESIS SUBMITTED IN PARTIAL F U L F I L L M E N T OF T H E REQUIREMENTS FOR T H E D E G R E E OF M A S T E R OF APPLIED SCIENCE in T H E 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 tofthe required standard  T H E UNIVERSITY OF BRITISH C O L U M B I A  October 1995  © R o y J.M. vanLier, 1995  In presenting this thesis in partial fulfillment of the requirements for an advanced degree at the University of British Columbia, I agree that the Library shall make it freely available for reference and study. I further agree that permission for extensive copying of this thesis for scholarly purposes may be granted by the head of my department or by his or her representatives.  It is understood that copying or  publication of this thesis for financial gain shall not be allowed without my written permission.  Department of M e t a l s  and M a t e r i a l s  The University of British Columbia Vancouver, B.C., Canada Date: O c t o b e r 27, 1995  Engineering  ii  Abstract  ABSTRACT  The INCO Ltd. Copper Refinery Electrowinning Department (CRED) in Copper Cliff, Ontario, processes a Cu, Ni, Co, Fe, As, S and precious metals containing residue produced by the INCO Pressure Carbonyl (IPC) plant.  This residue is treated batchwise in two successive leaching  stages. The non-oxidative First-Stage Leach (FSL) process extracts the vast majority of the Ni, Co, Fe and As values. The Second-Stage Leach (SSL) process comprises the aqueous pressure oxidation of a predominantly Cu S feed under acid-deficient conditions. 2  The basic copper  sulfate (BCS) precipitated in this stage is subsequently redissolved with spent electrolyte to separate the copper values from the final residue rich in precious and platinum group metals. The present study is part of a joint INCO-UBC research effort aiming at the optimization of both leaching stages. The purpose of this project is to provide a better understanding of the first-stage chemistry in general and of the behavior of arsenic in particular, since elevated concentrations of this element are detrimental to the efficiency of the second-stage system. The test work shows that the FSL process is insensitive to physicomechanical parameters such as agitation rate and pulp density, as long as the "solids off-bottom condition" is satisfied.  In  contrast, the effect of temperature and reactor residence time on impurity extraction is pronounced, especially in the case of arsenic. On the basis of thermodynamics and experimental observations made in the course of the investigations, a tentative reaction model is proposed to explain reprecipitation of this element (and nickel) under copper-depleted first-stage conditions.  The presence of organic electrowinning additives in solution does not influence the Ni, Co, Fe and As recoveries achieved in the first stage. The iron and, to a lesser extent, nickel extraction depend on the availability of acid in solution, whereas cobalt is leached principally by metathesis reactions.  Two scenarios for the leaching of arsenic are postulated, yet neither one can be  confirmed; the concentration of arsenic is either below the detection limit of the equipment (XRD) or is obscured by the presence of other elements (SEM).  Arsenic redissolution from a copper-depleted FSL batch cannot be accomplished by the introduction of pregnant copper electrolyte to the first-stage flash tank under the conditions of the laboratory simulation. In order to mitigate arsenic related second-stage leaching difficulties it is recommended that a potential probe be installed between the pressure let-down vessel and the filter feed tank. The on-line potential readings would allow timely changes in the make-up to avoid copper depletion of subsequent batches.  Abstract  iii  Simplified calculations have shown that the use of compressed air instead of "push steam" for discharging of first-stage reactors leads to an unacceptable increase in the copper level in solution.  Table of Contents  iv  TABLE OF CONTENTS  Abstract  ii  Table o f Contents  iv  List o f Figures  vii  List o f Tables  ix  List o f Abbreviations  x  Acknowledgments  xii  Motto and Dedication  xiii  Chapter 1 I N T R O D U C T I O N  1  Chapter 2 I N D U S T R I A L N I C K E L - C O P P E R M A T T E L E A C H I N G P R A C T I C E  4  2.1  Introduction  4  2.2  Chloride Lixiviants  5  2.2.1  Falconbridge technology  6  2.2.1.1  Matte Leach Process  6  2.2.1.2  Chlorine Leach Process  8  2.2.2 2.3  Eramet technology  11  Sulfate Lixiviants  13  2.3.1  A M A X technology  14  2.3.2  I N C O technology  15  2.3.3  Outokumpu technology  21  2.3.3.1  Finnish operations  21  2.3.3.2  Zimbabwean operations  24  2.3.4  Sherritt technology  24  2.3.4.1  Introduction  24  2.3.4.2  Impala Platinum  25  2.3.4.3  Rustenburg Base Metal Refiners  28  2.3.4.4  Western, Barplats and Northam Platinum  33  Table of Contents 2.3.5 2.3  v Mansfeld Kombinat technology  35  Nitrate Lixiviants  35  Chapter 3 D E V E L O P M E N T OF POURBAIX DIAGRAMS FOR T H E SYSTEMS ARSENICSULFUR-WATER AND NICKEL-SULFUR-WATER  37  3.1  Introduction  37  3.2  Sulfur-Water Diagrams  37  3.3  Arsenic-Sulfur-Water Diagrams  39  3.3.1  Ambient temperature  40  3.3.2  Elevated temperature  41  3.4  Nickel-Sulfur-Water Diagram  44  Chapter 4 E X P E R I M E N T A L W O R K  45  4.1  Sampling Program  45  4.2  Leaching Program  46  4.2.1  Scope  46  4.2.2  Batch make-up  47  4.2.3  Electrolyte preparation  48  4.2.4  Equipment  49  4.2.4.1  Autoclaves  49  4.2.4.2  Vacuum desiccators  50  4.2.5  Autoclave leaching procedure  51  4.2.6  Releaching procedures  53  4.2.6.1  Releaching with sodium hydroxide  53  4.2.6.2  Releaching with pregnant electrolyte  54  4.2.7 4.3  Hydrogen sulfide analyses  55  Sample and Waste Management  55  Chapter 5 RESULTS A N D DISCUSSION  '....57  5.1  Introduction  57  5.2  Sample Composition  58  5.3  Effect of Agitation and Pulp Density  59  5.3.1  Agitation series  59  5.3.2  Pulp Density series  62  5.4  Effect of Temperature and Residence Time 5.4.1  Kinetic studies on IPC residue with a high Cu : S  63 tot  ratio  63  Table of Contents 5.4.2 5.5  vi Kinetic studies on IPC residue with a low Cu : S  tot  ratio  68  Effect of Electrolyte Composition  69  5.5.1  Copper series  69  5.5.2  Acid series  70  5.5.3  Additives series  72  5.6  Reaction Model for First-Stage Leaching under Reducing Conditions  73  5.7  Behavior of Arsenic  74  5.7.1  Arsenic series  74  5.7.2  Releaching with sodium hydroxide  77  5.7.3  Releaching with pregnant electrolyte  78  5.8  Air Discharge Series  79  Chapter 6 CONCLUSIONS A N D RECOMMENDATIONS  81  References and Bibliography  84  Appendix A T H E R M O C H E M I C A L C A L C U L A T I O N S  91  Appendix B EXPERIMENT DESIGN C A L C U L A T I O N S  101  Appendix C EXPERIMENT WORKSHEETS  102  Appendix D AIR DISCHARGE C A L C U L A T I O N S  161  List of Figures  vii_  LIST O F F I G U R E S  Figure 2-1.  Schematic flowsheet of the Falconbridge Matte Leach (FML) process  6  Figure 2-2.  Schematic flowsheet of the Falconbridge Chlorine Leach (FCL) process  9  Figure 2-3.  Simplified flowsheet of the proposed and piloted SLN Refinery process  12  Figure 2-4.  Simplified flowsheet of AMAX's Port Nickel Refinery  14  Figure 2-5.  Simplified flowsheet of iNCO's CRED plant  17  Figure 2-6.  Schematic flowsheet of Outokumpu's Harjavalta Works  21  Figure 2-7.  Simplified flowsheet of the Impala Platinum nickel-copper refinery  26  Figure 2-8.  Simplified flowsheet of the Rustenburg Base Metal Refiners (RBMR) process  29  Figure 2-9.  Conceptual flowsheet of the Western, Barplats and Northam Platinum nickel-copper refineries  Figure 3-1.  "Standard" sulfur-water diagram at 25°C with 1 atm gas pressure and unit activity of all solutes  Figure 3-2.  40  Extended arsenic-sulfur-water diagram at 160°C with 1 atm gas pressure and 1 molal activity of aqueous arsenic and sulfur species  Figure 3-5.  39  Extended arsenic-sulfur-water diagram at 25°C with 1 atm gas pressure and 1 molal activity of aqueous arsenic and sulfur species  Figure 3-4.  38  "Extended" sulfur-water diagram at 25°C with 1 atm gas pressure and unit activity of all solutes  Figure 3-3.  34  43  Extended nickel-sulfur-water diagram at 25°C with 1 atm gas pressure and 1 molal activity of aqueous nickel and sulfur species  44  List of Figures  viii  Figure 4-1.  Sample taking from the sample point of the IPC thickener underflow line  45  Figure 4-2.  Autoclave set-up for the leaching program at U B C  50  Figure 4-3.  Vacuum desiccator set-up for the leaching program at U B C  51  Figure 5-1.  Variability in the grades of the principal impurity elements in IPC residue samples of May 9-24, 1994  59  Figure 5-2.  Effect of agitation rate on the extraction of cobalt, nickel, iron and arsenic  60  Figure 5-3.  S E M photograph showing "honeycomb" structures of different porosity  61  Figure 5-4.  Effect of pulp density on the extraction of cobalt, nickel, iron and arsenic  62  Figure 5-5.  Fraction of metal extracted from IPC residues with an average Cu : S of 4.3, as a function of leaching time  Figure 5-6.  Stacked X R D patterns showing mineralogical changes in the leach residue over  Fraction of metal extracted from IPC residue with a Cu : S function of leaching time  Figure 5-8.  67  tot  mass ratio of 3.2, as a 69  Effect of electrolyte copper concentration on the extraction of cobalt, nickel, iron and arsenic  Figure 5-9.  mass ratio 65  time at 160°C Figure 5-7.  tot  70  Effect of electrolyte sulfuric acid concentration on the extraction of cobalt, nickel, iron and arsenic  71  Figure 5-10. Extraction of cobalt, nickel and iron achievd in the arsenic series, as a function of leaching time  75  Figure 5-11. Arsenic precipitation and end-of-leach slurry potentials as a function of leaching time  75  List of Tables  LIST O F T A B L E S  Table 4-1.  Summary of First-Stage Leach (FSL) parameters investigated in the leaching program  Table 5-1.  Minimum, maximum and average grades of the main constituents of IPC residue samples taken from May 9 to 24, 1994  Table 5-2.  47  58  Effect of organic electrowinning additives on the extraction of cobalt, nickel, iron and arsenic  73  Table 5-3.  Effect of residue type and alkalinity on the re-extraction of arsenic  77  Table A-1.  Thermodynamic data of water species  91  Table A-2.  Thermodynamic data of sulfur species  91  Table A-3.  Thermodynamic data of arsenic species  92  Table A-4.  Thermodynamic data of nickel species  92  Table A-5.  Thermodynamic data used in the calculation of the standard entropy of the arsenyl ion  Table A-6.  93  Thermodynamic data used in the calculation of the standard entropy of the di-orthoarsenite and ortho-arsenite ions  94  List of Abbreviations  x LIST O F A B B R E V I A T I O N S  AAS  Ammoniacal Ammonium Sulfate  AMAX  American Metals ClimAX  BCL  Bamangwato Concession Limited  BCS  Basic Copper Sulfate  BNC  Basic Nickel Carbonate  B(Ni,Co)C  mixture of Basic Nickel and CObalt Carbonates  CCCR  Copper Cliff Copper Refinery  CCNR  Copper Cliff Nickel Refinery  CRED  Copper Refinery Electrowinning Department  D2EHPA  Di-2-Ethyl-Hexyl Phosphoric Acid  DSA  Dimensionally Stable Anode  EDX  Energy-Dispersive X-ray  EW  ElectroWinning  FCL  Falconbridge Chlorine Leach  FML  Falconbridge Matte Leach  FSL  First-Stage Leach  ICP  Inductively Coupled Plasma  INCO  The International Nickel COmpany of Canada  IPC  I N C O Pressure Carbonyl  IPL  International Plasma Laboratory  IUPAC  International Union of Pure and Applied Chemistry  IX  Ion eXchange  JRGRL  J . Roy Gordon Research Laboratory  MPP  Matte Processing Plant  xi  List of Abbreviations NBS  National Bureau of Standards  NRC  Nickel Refinery Converter  O/F  OverFlow  ORP  Oxidation-Reduction Potential  PGM's  Platinum Group Metals  P(G)M's  mixture of Precious and Platinum Group Metals  PHT  Product Holding Tank  PM's  Precious Metals  RBMR  Rustenburg Base Metal Refiners  SCE  Saturated Calomel Electrode  SEM  Scanning Electron Microscopy  SHE  Standard Hydrogen Electrode  SLN  Societe Le Nickel  SS  Stainless Steel  SSL  Second-Stage Leach  SX  Solvent extraction  TBP  Tri-Butyl Phosphate  TIOA  Tri-Iso-Octyl Amine  TOL  Total Oxidative Leach  UBC  The University of British Columbia  U/F  UnderFlow  UN  United Nations  USGS  United States Geological Survey  XRD  X-Ray Diffraction  Acknowledgments  xii  ACKNOWLEDGMENTS  The present research project would not have been possible without the help of a great number of dedicated people.  Most of all, I would like to thank Dr. David Dreisinger for giving me the  opportunity to do research in Canada. I feel privileged to have this Canadian experience and hope that more students will benefit from the "Vancouver-Delft connection" in the future. I am greatly indebted to the people at INCO for sponsoring both the metallurgical test work and the chemical analyses. I appreciate the thought-provoking progress meetings held with Lola Skelton and other staff of the Copper Refinery in Copper Cliff, and Dr. Eberhard Krause of the J. Roy Gordon Research Laboratory in Mississauga. My gratitude extends to faculty, colleagues and technicians at U B C .  Dr. Ernest Peters,  Dr. David Dreisinger, Ishwinder Grewal and Benjamin Saito all deserve many thanks for their valuable contributions to the understanding and improvement of the leaching operation at the C R E D plant.  I am thankful to summer student Mark Szeto for his help with the releaching  experiments.  The help of Ross McLeod, Carl Ng and Serge Milaire with the construction and  maintenance of the laboratory autoclave is also gratefully acknowledged.  Finally, many thanks  are due Mary Mager for her assistance in the S E M investigations, and Dr. Mati Raudsepp for his advice on the X R D analyses. Last but not least, I am most indebted to my wife Laurence as well as my family back home in Europe for their unconditional support and encouragement.  Motto and Dedication  xiii  MOTTO "Engineers are the pride of the nation "  DEDICATION To my lovely wife Laurence To my parents and my sister  Chapter 1  Introduction  1  CHAPTER 1  INTRODUCTION  The title of this thesis clearly shows both the company- and process-specific nature of its topic. In order to place the present investigations within the context of previous joint INCO-UBC research efforts, it is necessary to first review the metallurgical difficulties encountered in the two-stage leaching operation of INCO's Copper Refinery Electrowinning Department (CRED). In the oxidative Second-Stage Leach (SSL) process practiced at the C R E D plant of INCO's Copper Refinery at Copper Cliff, Ontario, the intermediate chalcocite (Cu S) residue from the 2  metathetic First-Stage Leach (FSL) process is batch pressure leached under acid-deficient conditions to produce a high-strength copper electrolyte for electrowinning and a residue rich in precious and platinum group metals for further upgrading.  The Cu S-containing residue is 2  converted to covellite (CuS), C u S 0 , and the basic copper sulfate antlerite (CuS0 -2Cu(OH) ) 4  4  2  within 1 hour. Subsequently, CuS is oxidized to C u S 0 in the presence of the basic copper 4  sulfate (BCS) precipitate. This takes 4-5 hours for completion under normal circumstances, but may take as long as 20 hours under so-called "slow cook" conditions. Plant experience shows that 10-20% of the SSL batches require those throughput-limiting extended reactor residence times.  In recent years, several collaborative studies carried out by C R E D personnel, research staff at INCO's J. Roy Gordon Research Laboratory (JRGRL), and faculty and students of UBC's Department of Metals and Materials Engineering, have focused on specific chemical and physicomechanical SSL process parameters in order to identify and mitigate the "slow cook" phenomena. The earliest laboratory work [1,2] concentrated on the verification of a mathematical screening model for the prediction of the SSL reaction sequence under various operating conditions. The results indicated that chemically controlled kinetics cannot be responsible for the periodic upsets experienced in the C R E D operation.  A gas-liquid mass transfer rate  controlling mechanism was proposed instead. Subsequent research [3] established a link between the feed composition, the apparent slurry viscosity and the copper extraction rate. It became evident that long second-stage leaching conditions are chiefly associated with two types of troublesome feeds: the high-(Cu : S) feed and the high-arsenic feed.  Chapter 1  Introduction  2  The most recent investigations [4,5] elaborated on copper extractions achieved from BCS precipitated at different leaching temperatures and with different levels of arsenic in the batch. They have shown that careful temperature control is critical to the SSL process, especially when high-arsenic feeds are being leached.  High-(Cu : S) feeds to the C R E D plant are caused by excessive sulfur removal prior to INCO's carbonylation process at the Nickel Refinery. The concentration  of BCS in the second-stage autoclave depends mainly on the feed  solids content and on the (Cu : S )  total  ratio in second-stage batch make-up.  The latter is  obviously affected by both the feed Cu : S ratio and the quantity of sulfuric acid - as spent copper electrolyte - utilized. Batches of high-(Cu : S) feeds can be leached more rapidly in the plant by increasing the acid content during second-stage batch make-up. This lowers the (Cu : S )  total  ratio, postpones the  onset of BCS formation and reduces the final concentration of antlerite. Injection of acid into the autoclave at the beginning of CuS leaching has a similar effect: it lowers the pulp density by partially redissolving the BCS already present.  Regardless of the mode of acid addition, the  concentrations and viscosities of the BCS slurries decrease and the covellite leaching rates improve. Unfortunately, the introduction of greater quantities of spent electrolyte to the batch by either method also promotes the undesirable formation of elemental sulfur.  In addition, it  increases the risk of crystallization of copper sulfate hydrates upon cooling of the SSL slurry during discharge. Furthermore, the addition of extra arsenic-containing spent copper electrolyte increases the amount of this unwanted element in the SSL system. Therefore, the only acceptable way to overcome the problems related to high-(Cu : S) feeds is to operate the SSL process at a lower pulp density, which ultimately limits the capacity of the leaching plant.  Sources of high-arsenic feeds are either residues produced under copper-deficient F S L conditions, or incompletely leached F S L residues.  Copper depletion, of course, results in a  change in the redox potential of the FSL system, causing reprecipitation of any dissolved arsenic, presumably as a base metal arsenide or arsenic sulfide. The morphology of BCS precipitated from high-arsenic feeds and common feeds is markedly dissimilar. Relatively large, compact, well-defined platelets are formed under normal circumstances.  Clusters of small, irregularly shaped or submicron acicular (needle-like)  particulates occur in the presence of high levels of arsenic in the SSL system. The formation of such aggregates is accompanied by an increase in slurry viscosity and simultaneous drop in copper extraction.  Chapter 1  3  Introduction  Interestingly, a higher degree of acicularity is noticed for the copper-depleted source than for the incompletely leached source. Furthermore, arsenic contained in the spent copper electrolyte has the same negative impact on the SSL process as arsenic present in FSL residues of the copperdepleted type.  It is believed that the modification of antlerite crystal shapes results from a  heterogeneous nucleation effect with some arsenic species acting as a BCS nucleation-catalyzing substrate during the early stages of second-stage leaching. Such a substrate, likely composed of the basic copper arsenate olivenite (Cu As0 (OH)), decreases the critical supersaturation 2  4  requirements for nucleation and increases heterogeneous nucleation rates at the expense of crystal growth rates. Hence, it is crucial the total arsenic level of the SSL batch be minimized. SSL temperatures higher than 115°C enhance problematic BCS characteristics via higher supersaturation levels from increased initial rates of copper sulfide leaching, by decreased nucleation activation energies and through decreased precipitate solubility. The combined effect of high feed arsenic content and higher leach temperature is particularly detrimental to copper extraction kinetics. Methods to alleviate the SSL process difficulties related to high-arsenic feeds are (1) to reduce the target FSL solids concentration in batch make-up, (2) to improve temperature control, and (3) to increase the agitation efficiency in order to improve gas-liquid mass transfer. Most of all, however, the extraction of arsenic achieved during first-stage leaching should be maximized to mitigate the deleterious role of the element in second-stage leaching. It is this insight that has instigated the current INCO-UBC investigations. Thus, the principal aim of the present project is to provide a better understanding of the chemistry and kinetics of the FSL process, particularly with respect to arsenic, in order to enhance the metallurgical performance in the SSL process.  In addition, it is hoped that the  experimental results will aid in the development of a superior control strategy for the C R E D leaching operation.  This thesis supersedes any views expressed and results reported previously [6,7]. It is organized in six chapters.  Chapter 2 is an inventory of contemporary industrial nickel-copper matte  leaching technology.  The chapter includes a general overview of INCO's operations in the  Sudbury Basin as well as a more detailed description of the C R E D plant. In Chapter 3 selected potential-pH diagrams  for the  systems arsenic-sulfur-water  and nickel-sulfur-water are  developed. Chapter 4 describes the experimental program, the results of which are summarized and discussed in Chapter 5. Finally, conclusions and recommendations for future research follow in Chapter 6.  Chapter 2  4  Industrial Nickel-Copper Matte Leaching Practice CHAPTER 2  INDUSTRIAL  2.1  NICKEL-COPPER  MATTE LEACHING  PRACTICE  Introduction  In this chapter, matte is defined as a generally sulfur-deficient pyrometallurgical product, essentially consisting of various nickel, copper and iron sulfides, a metallic phase with native metals and alloys, and both "simple" and spinel-type oxides. The actual concentrations of nickel, copper, cobalt, iron, sulfur, precious metals (PM's), platinum group metals (PGM's) and other elements can greatly vary, although the mineralogical composition of most mattes is quite similar. The alloys are important to the metallurgical industry since (1) during smelting they often act as scavengers for P(G)M's, and (2) they are magnetic, allowing relatively easy magnetic separation, and hence, concentration of the P(G)M values, after slow cooling of the matte. At present INCO and Rustenburg Platinum Mines are the only two companies in the western world to process iron-nickel-copper alloys separately; whole matte oxidative pressure leaching and, to a lesser extent, matte  electrolysis,  are practiced by the  majority of modern integrated  hydrometallurgical matte treatment plants in the world. The traditional, mainly pyrometallurgical route of matte smelting, slow cooling and separation, followed by roasting and reduction melting of the nickel sulfides and fire refining of the copper sulfides to produce crude metallic anodes for conventional electrorefining circuits of both metals, is presumably still employed at the Russian Monchegorsk and Norilsk plants [8].  Considering its limited bearing on the present leaching studies, matte electrorefining is not discussed in detail here. In such an electrolysis process matte is anodically dissolved, leaving a voluminous layer of elemental sulfur which acts as an effective collector for P(G)M's as well as selenium and tellurium. Depending on the composition of the anode, either copper or nickel may be electrodeposited at the cathode.  The production of copper powder from corroding bagged  matte anodes in a all-sulfate medium was formerly practiced by Engelhard Industries in New Jersey [9]. A similar process for a mixed sulfate-chloride electrolyte was patented in the Soviet Union [10].  Nickel cathodes were previously produced by Shimura Kako in Japan [11].  Currently, the production of nickel cathodes from electrodissolving matte anodes in a sulfatechloride(-borate) electrolyte with divided cells is carried out at INCO's Thompson Refinery [12], the Jinchuan Non-ferrous Metals Complex in China [13] and Sumitomo's Niihama Refinery in Japan [14].  Chapter 2  5  Industrial Nickel-Copper Matte Leaching Practice  In view of the heterogeneous composition of matte, it is not surprising that hydrometallurgical process flowsheets for the treatment of nickel-copper mattes are among the most complex and diverse in the metallurgical industry. Only in South Africa, and likely soon in Zimbabwe, are PGM's the primary matte products and base metals and other elements by-products. Eramet of France is the only company in the western world to solely process a virtually copper-free nickel matte produced from laterite ore.  The economic distinction between "nickel-copper mattes",  "precious metals mattes" and "nickel mattes" is not valid in this chapter, since all of the mattes are subjected to the same kind of leaching processes.  Nickel and cobalt may be marketed as sheared cathode slabs or special cathode shapes, as mixed precipitates or crystalline salts, or as powders, pellets and briquettes. Copper is usually sold as cathodes. Few refineries in the world produce separate P(G)M products, so that PM's and PGM's are often traded as concentrates. Sulfur is either recovered as S°, N a S 0 or ( N H ) S 0 for sale 2  to the chemical industry, or rejected as CaS0 . 4  4  4  2  4  Most leaching and electrorefining plants also  make selenium and tellurium products. To satisfy changing market demands, matte processing has been a field of continuous hydrometallurgical innovation. This chapter, therefore, describes the evolution of commercial nickel-copper matte leaching processes as part of an inventory of available technology.  It is  meant to complement and update parts of the established works [8,15]. Emphasis is put on the leaching chemistry in chloride, sulfate and nitrate media; the individual metal recovery processes further downstream are not addressed in great depth.  Since accurate information about plant  practice in the former Soviet Union is generally unavailable, the following review focuses on the industrial "know-how" in the western world.  2.2  Chloride Lixiviants  Chloride solutions are highly suitable media for matte processing.  By virtue of a lower  electrolyte resistivity, they offer the advantages of a lower cell voltage and a higher current density in nickel electrowinning when compared to predominantly sulfate media. Copper can exist in two valency states in acidic chloride solutions and thus serve as an "electron carrier" for the electrochemical matte leaching reactions.  Finally, the economically important removal of  cobalt from solution is greatly facilitated when the substantial chloride concentrations introduced through oxidative hydrolysis of this metal with C l - N i C 0 can be tolerated and accommodated 2  by the nickel electrolyte.  3  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  6  Unfortunately, the aggressive nature of chloride lixiviants and fear of process control difficulties have limited the use of hydrochloric acid and chlorine gas in the metallurgical industry. To date there are only two companies that successfully operate all-chloride nickel and cobalt flowsheets: Falconbridge of Norway and Eramet of France.  2.2.1  Falconbridge technology  2.2.1.1  Matte Leach P r o c e s s  The Falconbridge Matte Leach (FML) process was engineered for and piloted at the Falconbridge Nikkelverk in Kristiansand, Norway.  It was implemented by the late 1960's to bridge the  transition of the refinery's old modified Hybinette flowsheet to the chlorine leach process practiced today.  At the time the refinery processed nickel-copper convertor matte from the  Falconbridge smelter in Sudbury, with typically 48% nickel, 27% copper, 22% sulfur, about 1% each of cobalt and iron, as well as impurities such as arsenic, lead and selenium.  H0 2  Figure 2-1.  Schematic flowsheet of the Falconbridge Matte Leach (FML) process [16].  Chapter 2  7  Industrial Nickel-Copper Matte Leaching Practice  Figure 2-1 is a schematic flowsheet of the F M L process [16]. The requirements of the pilot plant were drawn as a 98% minus 325 mesh fraction. Non-oxidative leaching of the matte with strong hydrochloric acid at 70°C in a four-stage co-current cascade of mechanically agitated rubberlined tanks selectively dissolved 98% of the nickel, leaving copper and P(G)M's as an insoluble sulfide residue. The total residence time in the leaching circuit was about 12 hours. The Cu S residue obtained after filtration of the leach slurry served as the feed material to 2  the copper section of the refinery's older modified Hybinette flowsheet [17].  In this way the  P(G)M's were more directly concentrated as solids, resulting from the copper leach, instead of as nickel anode slimes. Gas formed during the nickel leach consisted of a mixture of H and H S . It was cooled, 2  scrubbed free of HC1 and combusted in a waste heat boiler.  2  The combustion gases were  combined with the roaster off gases and treated for recovery of liquid S 0 . 2  The pregnant nickel solution, containing approximately 120 g/L nickel and 160 g/L HC1 as well as about 2 g/L each of iron, cobalt and copper, was purified in three stages. First, oxygen was passed through the solution to oxidize dissolved H S to S° and ferrous iron to its ferric state. 2  After filtration for removal of the sulfur, the solution was cooled to room temperature in preparation for solvent extraction. Iron(ITf) was then removed with the solvating extractant tributyl phosphate (TBP). Finally, treatment of the iron-free raffinate with solvating tri-iso-octyl amine (TIOA), followed by selective stripping, removed cobalt and copper. A l l stripping was carried out with water. Lead was the only remaining impurity after solvent extraction treatment.  Since it was  highly soluble under the crystallizer conditions - through complexation as chloride anions - and caused no serious contamination of the crystallizer product, it was removed from the acid regeneration system in a small ion exchange column. In the continuous crystallizer HC1 gas was injected into the purified pregnant nickel solution to "salt out" crystals of N i C l - 4 H 0 . 2  2  Following centrifuging, mother liquor with roughly 28 g/L  nickel and 330 g/L HC1 flowed into strong acid storage tanks, whereas the crystals discharged into a bin feeding a direct-fired rotary dryer. In the dryer about 3 of the 4 moles of water of crystallization were removed from the crystals.  Partly dehydrated nickel chloride was then fed to a high-temperature fluidized-bed  hydrolysis reactor ("converter"), in which rounded high-density particles of nickel oxide were formed according to the reaction: NiCl  2  + H0 2  NiO + 2 HCl  (2-1)  Chapter 2  8  Industrial Nickel-Copper Matte Leaching Practice  The NiO granules were continuously discharged from the bottom of the pyrohydrolyzer, cooled, and screened for recovery of minus 35 mesh material, which was recycled to serve as seed material for the vapor deposition hydrolysis reactions. The converter off gases, before passing through HC1 absorbers, passed directly through a cooling tower fed with purified pregnant solution.  The latter underwent  considerable  concentration in nickel in this operation, emerging as crystallizer feed. This enrichment resulted partly from evaporation and partly from dissolution of NiO fines which were carried over with the converter off gases. Unfortunately, it also led to the presence of some nickel oxide in the crystalline product. The cooled gases entered the HC1 adsorption system for recovery of the regenerated acid. Reduction of NiO to a high-purity granular nickel metal product, designated N I C K E L 98, was done in a rotary hydrogen reduction furnace.  2.2.1.2 Chlorine Leach P r o c e s s The modified Hybinette process was in use at the Kristiansand Nikkelverk until the late 1970's, when conversion to the Falconbridge Chlorine Leach (FCL) process was initiated.  Nickel  electrowinning from all-chloride electrolyte [18] has since replaced electrorefining of the metal in a chloride-sulfate electrolyte.  A blend of converter mattes, mainly from the Falconbridge  smelter in Sudbury and the B C L smelter in Botswana, is currently being processed. The feed to the refinery typically contains 40-45% nickel, 25-30% copper, 20-22% sulfur, 1-1.5% cobalt and 2-3% iron. Figure 2-2 is a schematic flowsheet of the F C L process [19]. The matte consists principally of heazlewoodite (Ni S ), chalcocite (Cu S) and an alloy with Ni : Cu = 7 : 3. It is wet-ground to a 3  2  2  particle size wholly finer than 80 mesh and advanced to the atmospheric leaching tanks. Chlorine from the nickel tankhouse is fed under the impeller and the temperature is held at the boiling point of the slurry. The main leaching reactions [20] are:  2 CuCl + CI2 —> 2 Ni S 3  2  + 2 CuCl  -» NiCl  2  2  Cu S 2  2  + 2 NiS + 2 CuCl  2  NiS + 2 CuCl  CuCl  -> NiCl  2  + 2 CuCl + S°  + S° -> 2 CuS  (2-2) (2-3)  (2-4)  (2-5)  Chapter 2  Figure 2-2.  Industrial Nickel-Copper Matte Leaching Practice  Schematic flowsheet of the Falconbridge Chlorine Leach (FCL) process [19].  9  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  10  The leaching process is based on control of the oxidation-reduction potential (ORP) to maintain a desired ratio of cupric to cuprous ion; at all times sufficient CuCl must be available to ensure immediate absorption of the chlorine as the nickel leaching intermediary C u C l . 2  Leach slurry at 110°C is pumped from the leach tanks through a steam preheater where its temperature is raised to about 130°C, and then to the first of two autoclaves in which the exchange reactions are sufficiently exothermic to raise the temperature to 140-145°C. Although in the autoclaves the principal reactions take place between the solution and unleached millerite (NiS) in the chlorine leach residue, any tendency for the ORP and dissolved copper to vary beyond set limits is automatically adjusted by the introduction of minor amounts of fresh matte. The slurry flows continuously into the second autoclave and is discharged via a cooling system to the first of two "cementation" tanks in tandem. The leaching chemistry of the autoclave train [19] is best described by: NiS + 2 CuCl -> NiCl  + Cu S  2  (2-6)  2  The reactions [19] between the leached slurry and the fresh matte introduced in the copper precipitation tanks are: Ni S 3  2  + S° + 2 CuCl -> NiCl  + NiS + Cu S  2  Ni + S° + 2 CuCl S°  -> NiCl  2  + 2 CuCl -> CuCl  2  (2-7)  2  + Cu S  (2-8)  2  + CuS  (2-9)  The formation of covellite (CuS) through reaction (2-5) occurs throughout the above circuit. The overall nickel extraction of the leaching-copper precipitation system is 90%.  The copper  concentration decreases from 50-70 g/L in the leach tanks to 0.5 g/L in the second copper precipitation vessel. After filtration, the resulting strong nickel chloride solution is purified and sent to the nickel tankhouse. Various nickel products are produced by electrowinning, including cathodes and "crowns". Anodes (DSA's) are fitted with a hood, a diaphragm bag, and a duct connecting the top of the hood to a manifold running alongside the cell for withdrawal of chlorine and anolyte by suction.  Purification of the pregnant nickel solution involves (1) precipitation of iron and arsenic with C l  2  and N i C 0 , (2) precipitation of gypsum through cooling, (3) solvent extraction of cobalt with 3  TIOA, and (4) precipitation of lead, manganese and other trace impurities with C l and N i C 0 , 2  3  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  after dilution of the nickel stream with anolyte.  11  The all-chloride pregnant cobalt solution  contains 80-90 g/L cobalt and is pumped to the cobalt tankhouse. The copper sulfide chlorine leach residue acts as a scavenger for P(G)M's. It is dead roasted in fluidized-bed roasters. H S0 . 2  4  Sulfur dioxide is recovered from the roaster off gases and marketed as  The calcine is leached in spent copper electrolyte and copper is electrodeposited from  the pregnant solution.  Using some of the idle F M L equipment, the copper leach residue is  subjected to lixiviation with HC1-C1 [21] to solubilize residual nickel and copper while leaving 2  the P(G)M's intact. This step also releases residual iron, which is precipitated and filtered off before the nickel-copper solution is sent back to leaching. The final crude P(G)M concentrate is further treated in the precious metals refinery.  2.2.2  Eramet technology  Eramet (formerly SLN) in Le Havre-Sandouville, France, developed and still operates the only other commercial all-chloride process for the integrated recovery of nickel and cobalt from matte. The matte, shipped from the company's Doniambo smelter in New Caledonia, contains virtually no copper. Since it has proven difficult to find detailed information about the present refinery practice, this section chiefly discusses the original pilot-plant campaign carried out at Minemet Recherche in the mid 1970's for the development of a hydrometallurgical flowsheet for the S L N Refinery [22]. Figure 2-3 is a simplified flowsheet of the originally piloted process. Matte was crushed, ground and fed to a fluidized-bed roaster. After roasting, the NiO calcine, with minor metallic nickel and nickel sulfide, was leached with strong (8A ) HC1 in a two-stage countercurrent leaching 7  circuit, kept at the boiling point of the slurry: NiO + 2 HCl -> NiCl  + H0  (2-10)  2  +H  (2-11)  2  + HS  (2-12)  2  2  Ni + 2 HCl -» NiCl NiS + 2 HCl -> NiCl  The residence time was 4 hours in each tank.  2  2  Chapter 2  Industrial Nickel-Copper  12  Matte Leaching Practice ground matte S02^  roasting  I  HCI  zn leaching  CI2  H2SO4 production  -residue  Fe oxidation  I  eluate  IX  eluate adjustment  |NiCl2  NiCOs Pbmixed hydroxides  Pb electrolysis NiCOs*^ precipitation  trace base metals HCI regeneration  Figure 2-3.  Fe SX  IX  I  Iraffinatel  ^  t  FeCIs  Co SX strip|solution Co EW  Ni EW  ^ NiC03 productionlZZf H2  evaporation  ibleed stream  ,  r  Co cathodes  Ni cathodes  Simplified flowsheet of the proposed and piloted SLN Refinery process (after [22]).  Following liquid-solid separation, the leach solution was advanced to an iron oxidation step, where at 75°C an upward flow of chlorine gas was bubbled through a packed column to ensure the formation of the extractable Fe(HI) species F e C l in the descending leach solution. Traces of 4  elemental sulfur, formed by the oxidation of residual dissolved H S , were then filtered off the 2  N i C l stream prior to further purification stages. 2  In two ion exchange columns in series chloride complexed cobalt, copper, iron and zinc were removed using the strong basic anion exchanger IRA 400. Evidently, the ORP of the chlorinetreated leach solution had to be carefully controlled to prevent degradation of the resin. The chloride level of the eluate was adjusted to 57V by the addition of some N i C l . The 2  solution was then concentrated to approximately 160-170 g/L nickel, 30 g/L iron and 20 g/L cobalt by evaporation, before solvent extraction of iron with Amberlite L A 2 and cobalt with TIOA. Cobalt was subsequently recovered by electrodeposition. the nickel solution from ion exchange.  The raffinate was mixed with  Chapter 2  Industrial Nickel-Copper  The pH of the pregnant N i C l  2  13  Matte Leaching Practice  solution was gradually increased with N i C 0  3  for purification.  First, lead was removed through selective electrowinning at pH 2. Secondly, mixed hydroxides of lead, chromium and aluminum were precipitated at pH 3.5.  The third purification step  comprised ion exchange with a small column to remove the last traces of base metals. Build-up of the remaining impurities - sodium, calcium, magnesium and sulfate - was prevented by thorough washing of the N i C 0 cake produced through addition of N a C 0 to a small stream of 3  2  3  catholyte. A l l impurity residues required further processing for re-extraction of entrained nickel.  Nickel was electrowon in diaphragm cells with air-mixing of the catholyte for improved mass transfer, and hence, deposition rates.  Chlorine was recovered at the anodes for production of  H C l for the leaching circuit. The major drawbacks of this rather cumbersome flowsheet were the necessity of production units for both hydrochloric and sulfuric acid.  In particular, the use of H  2  rendered the process  economics unfavorable. The finally adapted flowsheet [23] at the Eramet Refinery in 1978 is remarkably similar to the F C L process. In view of the low copper content of the Doniambo matte, chlorine leaching with FeCl as the leaching intermediary is employed instead. 3  Sulfur remains in its elemental  state in the leach residue. Iron(III) chloride is removed by solvent extraction with TBP. Part of it is recycled to the leaching stage, whereas the other portion is concentrated through evaporation and sold. Cobalt is extracted from the impure nickel chloride solution with TIOA and sold as solution for cobalt metal recovery elsewhere.  After a similar deleading electrolysis stage as  devised for the process described above, the N i C l  2  solution is passed through columns of  activated carbon to remove residual impurities. Nickel is electrowon onto titanium cathodes in electrolysis cells with graphite anodes.  2.3  Sulfate Lixiviants  Sulfuric acid is the most frequently used mineral acid in the hydrometallurgical processing of matte.  Its lack of selectivity and corrosive nature are largely offset by its low cost and wide  availability.  This section discusses the sulfuric acid based processes developed by A M A X ,  INCO, Outokumpu, Sherritt and the Mansfeld Kombinat.  Chapter 2 2.3.1  Industrial Nickel-Copper Matte Leaching Practice  14  A M A X technology  The A M A X Port Nickel Refinery in Louisiana was in production from 1974 to 1986. It was the only producer of pure nickel in the U S A during that period. Matte  Crushing Grinding  I  Primary ^ N i - C o Sep'n.  [Atmospheric Leach  Nickel Reduc t i o n  1st Pressure Leach  Nickel Product  Pentammine  1  Separation  T  I  I  Co-Cake Dissolution  1  2nd  !> ir~  Pressure Leach  i Flotation  i  Copper Tankhouse  T  Copper  Figure 2-4.  IX System  ]  T  Iron Tailings  Product  t  H S Scavenging  Cobalt Reduction  2  T  Cobalt  Product  Crystalli-  I  zation  Solids or Slurry Solution  Ammonium S u l f a t e Product  Simplified flowsheet of AMAX's Port Nickel Refinery [24].  Figure 2-4 is a simplified flowsheet of the Port Nickel plant [24].  The feed to the plant was  matte from Africa (BCL) and Australia, which was crushed, ground to 95% passing 200 mesh and blended. The average composition of the matte was 47% nickel, 30% copper, 20% sulfur, 1%> iron and 0.5% cobalt. The  nickel-copper matte was leached in a multi-stage  atmospheric and pressure leaching.  countercurrent circuit of  The leaching chemistry of such a circuit is covered in  sections 2.3.3 and 2.3.4. The residue, consisting almost entirely of iron oxide, was cleaned of unleached sulfides in flotation banks operating in closed circuit with the leaching plant.  No  reagents were needed to float the sulfide minerals due to the presence of the naturally hydrophobic layer of elemental sulfur that encapsulated them.  Copper was electrowon as cathodes.  Cobalt was precipitated from the impure N i S 0 solution  with ammoniacal ammonium peroxydisulfate:  4  Chapter 2 2 CoS0  4  Industrial Nickel-Copper Matte Leaching Practice + 6NH  + (NH ) S O  3  4  2  2  + 6 H0  s  2  -> 2 Co (OH)  3  15  + 4 (NH ) S0 4  2  (2-13)  4  Following liquid-solid separation, nickel powder was produced from the cobalt-free ammoniacal ammonium sulfate (AAS) solution in an adapted Sherritt hydrogen reduction process. powder was the basis for various sintered nickel products.  The  Cobalt metal was produced in a  modified Sherritt "pentammine" process, which involved (1) redissolution of the cobaltic hydroxide cake in a reducing H S 0 - C H O H lixiviant, (2) ammoniation, (3) ion exchange with 2  4  3  TP 207 resin to remove trace nickel and other divalent ions, and (4) hydrogen reduction of cobalt. Cobalt was sold as powder and as "cubes".  The nickel and cobalt reduction end solutions were combined and subjected to a H S scavenging 2  treatment in a pipe reactor at neutral pH and ambient temperature. fertilizer-grade ( N H ) S 0 4  2  4  was  After pressure filtration,  crystallized from the base metal depleted  solution by  evaporation.  2.3.2  INCO technology  INCO's matte processing facilities in Canada comprise (1) the aforementioned electrolytic Thompson Refinery, operated by the Manitoba Division, and (2) a vast metallurgical complex centered around the township of Copper Cliff in the Sudbury Basin, operated by the Ontario Division. Since the present thesis is an analysis of one of the intermediary leaching processes practiced at the Copper Refinery, a brief general overview of INCO's operations in the Sudbury Basin is first given. Five underground INCO mines currently produce the nickel-copper ores from the Canadian Shield. A l l run-of-mine ore is crushed, ground in a semi-autogenous grinding circuit, and floated in large Outokumpu flotation cells at the Clarabelle Mill to reject pyrrhotite and other gangue minerals. The mill has a capacity of 40,000 tons per day of ore and produces a bulk nickelcopper concentrate that proceeds to the smelter. At the smelter, a recent five-year, $600 million (Cdn.) S 0  2  abatement program has  replaced multi-hearth roasters, reverberatory furnaces and the majority of the convertors with two INCO designed flash furnaces. In addition, a new sulfuric acid plant and a liquid sulfur dioxide facility have been erected. The flash furnaces produce a matte assaying roughly 35% nickel and 35% copper, the balance being sulfur. In the separation section of the Matte Processing Plant (MPP), the matte is cast into molds and allowed to cool over a four-day period to promote the formation of large  Chapter 2 Ni S 3  2  Industrial Nickel-Copper Matte Leaching Practice  and Cu S crystals.  16  Since the matte is deficient in sulfur, a metallic phase containing a  2  magnetic iron-nickel-copper alloy is also formed. In the MPP the matte is crushed, rod-milled and passed over magnetic separators to remove the metallics, which report to the Copper Cliff Nickel Refinery (CCNR). magnetic fraction is ball-milled and pumped to flotation columns for C u S - N i S 2  3  2  The nonseparation.  Chalcocite reports to the froth phase, whereas heazlewoodite is collected at the bottom of the columns. After filtration, the copper sulfide (MK) filter cake is converted to blister copper and transported in hot metal rail cars to the Copper Cliff Copper Refinery (CCCR), where it is firerefined to anode quality, cast and finally electrorefined. Heazlewoodite is oxidized in fiuidizedbed roasters.  The resulting NiO calcine is either sold or further treated at the C C N R or the  Clydach Refinery in Wales. The C C N R is the western world's largest producer of nickel. It operates a high-pressure variation of Mond's carbonylation process that extracts the majority of the nickel and about a third of the iron, as gaseous Ni(CO) and Fe(CO) , respectively, from the metallic fraction mentioned earlier. 4  5  Carbonylation is carried out batchwise in three horizontal, cylindrical rotating reactors.  The  volatile nickel tetra- and iron pentacarbonyls are condensed and separated through distillation. The presence of Fe(CO) necessitates the production of a ferronickel product along with the 5  highly pure nickel. The C C N R consists of two individual plants, namely (1) the Nickel Refinery Converter (NRC) plant, where the alloy phase is melted in top-blown rotary converters together with secondaries of various sources, and granulated in preparation of carbonylation, and (2) the INCO Pressure Carbonyl (IPC) plant, where nickel is extracted and decomposed into pellets and powders of extreme purity. The IPC process leaves a porous, granular residue, which, after depressurizing and purging with inert gas, is discharged from the reactor, slurried with water and milled to typically 80% passing 325 mesh before pumping to the Copper Refinery Electrowinning Department (CRED) of the C C C R as a dilute (5% solids) slurry. The C R E D plant was commissioned in the early 1970's. Figure 2-5 is a simplified flowsheet of the current process [25,26]. The incoming IPC slurry is thickened to approximately 55% solids, at which pulp density the material is stored in three holding tanks. In this way, daily fluctuations in the composition of the residue are smoothed out and a fairly uniform composition of the feed over several batches is achieved. Some 50 tons of residue are processed daily.  Chapter 2  17  Industrial Nickel-Copper Matte Leaching Practice  IPC residue slurry impure pregnant steam  H2O  —1 —1  02  H2SO4  spent First Stage Leach  I 1  Second Stage Leach  I I I  BCS redissolution  repulp  7  P(G)M slurry to Port Colborne Se-Te removal  7  pure pregnant  Cu EW — • Cu cathodes  Fe-As removal  3 /  lime 1—  soda ash  O2  1  Ni-Co precipitation  L  1 Ni-Co redissolution  I  7 7  H2SO4 thickening  H2O  O/F  B(Ni,Co)C to Port Colborne  repulp  lime  i  effluent disposal tailings  1  Se-Te cake to Silver Refinery Figure 2-5.  U/F  Simplified flowsheet of INCO's C R E D plant (after [25,26]).  Chapter 2  18  Industrial Nickel-Copper Matte Leaching Practice  A constant volume tank suspended on load cells allows known quantities of thickened slurry to be advanced from the holding tanks into the first-stage batch make-up tank.  Here, desired  mixtures of spent and pregnant copper electrolytes are added to the thickened slurry, and steam is used to preheat the batch to 80°C.  The preheated batch is then gravity fed to one of three  mechanically agitated brick-lined first-stage autoclaves. In the first-stage autoclaves the 20% solids slurry of IPC residue, electrolytes and steam condensate together is maintained at 160°C and 80 psig "steam" pressure for one hour. The purpose of the First-Stage Leach (FSL) process is to dissolve nickel, cobalt, iron and arsenic, in order to separate these elements from selenium, tellurium and the P(G)M's, all of which remain in a substantially Cu S-containing residue. The total FSL cycle time including batch make-up, 2  heating, leaching and discharging is typically 3-3.5 hours. The mineralogical composition of the highly reduced IPC residue is mainly Cu S (1.76 < x < 2), x  pentlandite ( M S with M = Ni, Fe, Co), N i S , metallics (Cu, Ni), spinel-type oxides ( M F e 0 ) , 9  8  3  2  2  4  and oxidation products formed during transport and storage, such as cuprite (Cu 0), tenorite 2  (CuO) and bunsenite (NiO). It is black in appearance and has a density of about 5500 kg/m . 3  The first-stage leaching chemistry involves the dissolution of the oxide minerals: MO + H S0 2  MFe 0 2  4  + 4H S0 2  4  MS0  4  + H0  4  -» MS0  4  (2-14)  2  + Fe (S0 ) 2  4  + 4 H0  3  2  (2-15)  cementation reactions: M + CuS0  -> MS0  4  + Cu  4  (2-16)  and metathesis (or "double decomposition") reactions in which reductants such as cuprous ion are attributed an important role [27]: Cu + CuS0  -> Cu S0  4  MS 9  8  + 7 Cu S0 2  4  + 2CuS0  2  4  (2-17)  4  -> 9 MS0  4  + 8 Cu S 2  (2-18)  In order to obtain optimum extractions of impurities, the FSL process is conducted with excesses of both cupric ion and sulfuric acid, typically 1-5 g/L and 20-60 g/L more, respectively, than  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  19  required. Excess acid, of course, is needed to prevent hydrolysis of the extracted metal values. A copper clean-up step with NaHS was abandoned in 1988 because of operating difficulties, and any residual copper is now either lost to the tailings or reports to the nickel-cobalt product as an unwanted contaminant. Indeed, the residual copper concentration must be carefully controlled.  Insufficient  copper in solution at the end of the F S L produces low potentials causing "reprecipitation" of dissolved arsenic and enhanced corrosion rates due to the action of H S. 2  During discharge, too  high a level risks clogging of chokes and adjacent piping on the low-pressure side with elemental copper, formed through the disproportionation reaction: Cu S0 2  4  —> Cu +  (2-19)  CuS0  4  The F S L filtrate is subjected to a two-stage hot lime treatment. At a temperature of 100°C at 50 psig oxygen pressure and a pH increasing from about 3.2 in the first autoclave to 4.1 in the second one, a mixture of ferric hydroxide, copper and iron arsenates and gypsum is precipitated. After filtration the iron-arsenic cake is repulped with water and H S 0 2  4  (pH 2-2.2) to redissolve  co-precipitated nickel and cobalt. Another liquid-solid separation step then separates the nickelcobalt solution from the ferric arsenate residue, which is once more repulped and filtered before it is discarded. At this point the pregnant solution contains approximately 15 g/L each of nickel and cobalt. In two precipitation tanks in series kept at 75°C, the pH of the nickel-cobalt solution is adjusted to 7.7 and 8.8, respectively, by the addition of soda ash. The resulting mixture of basic nickel and cobalt carbonates (B(Ni,Co)C) is trucked to the Port Colborne Cobalt Refinery as thickener underflow slurry. At Port Colborne it is refined into electrolytic cobalt "rounds" and N i C 0 , which is trucked back to CRED for drying and packaging. 3  The F S L cake is recovered from the rotary vacuum filter and slurried to 25-30% solids with water and some spent copper electrolyte in the second-stage batch make-up tank. The batch is then charged into one of two mechanically agitated second-stage autoclaves. The Second-Stage Leach (SSL) process is a total oxidative leach (TOL) in an aciddeficient environment at 115°C and a total pressure of 150 psig . It yields a slurry of C u S 0 and 4  a residue of essentially di-basic copper sulfate (antlerite) and P(G)M's. The equilibrium between dissolved copper and antlerite buffers the solution pH between 2 and 3.  Simplifying the  chemistry for a chalcocite feed, the sequential reactions are: Cu S 2  + H S0 2  4  + j 0  2  -> CuS0  4  + CuS +  H0 2  (2-20)  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  Cu S 2  + j CuS0  + H0  4  2  + j 0  CuS + j CuS0  2  4  CuS + 2 0  •  20  2Cu(OH)  2  -> CuSO.4  2  (2-21)  (2-22)  Any metallic copper present in the FSL residue reacts according to: Cu + H S0 2  4  + j 0  -> CuS0  2  4  +  (2-23)  H0 2  while the formation of elemental sulfur is considered to be a minor side reaction under normal operating conditions: CuS + H S0 2  + j 0  4  -» CuS0  2  4  + S° +  (2-24)  H0 2  The brick-lined second-stage autoclaves are equipped with cooling coils which abstract the heat generated by the above exothermic reactions and serve as baffles at the same time. The retention time of a batch in the second-stage reactor is typically 4-6 hours, but may be as long as 20 hours under the "slow cook" conditions explained in Chapter 1. The second-stage slurry is mixed with spent electrolyte from the copper tank house to redissolve antlerite according to: CuS0  4  • 2Cu(OH)  2  + 2 H S0 2  4  -> 3 CuS0  4  + 4  H0 2  (2-25)  Perlite filter-aid assisted pressure filtration then separates the impure pregnant copper electrolyte from a high-grade P(G)M residue which, after repulping, is sent for further concentration at Port Colborne. After heating of the pregnant electrolyte to 95°C in a heat exchanger, it is pumped upward through a bed of copper wire choppings in a cementation reactor to remove selenium and tellurium from solution. Unfavorable kinetics necessitate a retention time of 12-14 hours in four consecutive plug-flow aging towers to drive the copper selenide and telluride formation reactions to completion. The purified C u S 0 solution is advanced to a conventional electrowinning process with 4  antimonial lead anodes and titanium blank sheets. The crude selenium-tellurium concentrate is routed to the C C C R "Silver Refinery" and integrated with the treatment of anode mud from the copper electrorefining process. The "Silver Refinery" produces metallic silver and gold, separate impure selenium and tellurium cakes and a PGM-rich residue for further upgrading at Port Colborne. The PGM's are finally separated at the Acton Refinery in England.  Chapter 2 2.3.3  Industrial Nickel-Copper Matte Leaching Practice  21  Outokumpu technology  2.3.3.1 Finnish operations Outokumpu's Harjavalta Works in Finland process a high-grade matte largely produced from concentrates and secondaries from the company's own mills and refineries. The matte is blown to a very low sulfur content (^8%) prior to entering the leaching plant. Historically, Outokumpu has employed atmospheric pressure leaching of the matte to extract and separate its nickel and copper values [8].  Apart from the obvious physicomechanical process  parameters, the maximum degree of dissolution of nickel and copper attainable in such a leaching process depends on the Ni : Cu ratio and sulfur content of the matte [28]. With ever decreasing Ni : Cu ratios in the ore and matte, the nickel extraction achieved with the original three-stage atmospheric leaching system became insufficient and resulted in unacceptably high circulating loads of nickel-copper residue to the smelter.  The addition of a selective oxygen pressure  leaching stage became therefore necessary. A 70 m five-compartment autoclave was added to 3  the Harjavalta operations in 1981 [28], followed by a seven-compartment reactor in 1986 [29].  Figure 2-6.  Schematic flowsheet of Outokumpu's Harjavalta Works [29].  Figure 2-6 is a schematic flowsheet of the Harjavalta Works [29].  Granulated matte is wet-  ground to 90% minus 270 mesh and proceeds to the countercurrent leaching circuit.  Chapter 2  22  Industrial Nickel-Copper Matte Leaching Practice  Until recently, 130 m  atmospheric leaching vessels were used in the atmospheric section.  3  Agitation was carried out by blowing air into the reactors through their conical bottoms, while oxygen at the same time acted as an oxidizer. Apparently, new in-house developed oxygenblown atmospheric reactors are now being utilized, which has improved the tolerance to iron and arsenic in the Outokumpu process.  The new reactors were installed as part of the latest  expansion at Harjavalta [30]. In the first-stage atmospheric leach, metallic nickel, present in the freshly introduced matte, cements copper from solution: Ni + CuS0  -> NiS0  4  + Cu  4  (2-26)  leaving an impure N i S 0 solution which is pumped to the purification section. 4  Only when excess metallic nickel is depleted through oxidation: Ni + H S0 2  4  + j 0  2  NiS0  + H0  4  (2-27)  2  the release of nickel into solution in the successive atmospheric leaching stages becomes dictated by the - kinetically also fast - solid state transformation of heazlewoodite to millerite: Ni S 3  2  + H S0 2  + j 0  4  2  -> NiS0  + 2 NiS + H 0  4  (2-28)  2  Insufficient atmospheric oxidation leaves unreacted heazlewoodite in the leach residue, whereas excessive oxidation leads to the unwanted formation of polydymite (Ni S ): 3  + j 0  4 NiS + H S0 2  4  2  -> NiS0  + Ni S  4  3  4  + H0  4  (2-29)  2  A millerite residue is preferred, since the metathetic leaching of NiS with C u S 0 to N i S 0 and 4  4  digenite (Cu S ) according to: 9  5  6 NiS + 9 CuS0  4  + 4 H0 2  -> 6 NiS0  + Cu S  4  9  is kinetically much faster than the reactions of N i S 3  prevailing autoclave conditions:  2  + 4 H S0  5  2  and N i S 3  4  4  (2-30)  with copper sulfate at the  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  8Ni S 3  5 Ni S  2  3  4  + 27 CuS0  4  + 4 H0  -> 24 NiS0  + 28 H 0  -> 24 NiS0  2  + 45 CuS0  4  + 3 Cu S  4  2  9  + 4 H S0  5  2  + 5 Cu S  4  9  23 (2-31)  4  + 28 H S0  5  2  4  (2-32)  To enhance the conversion of any residual heazlewoodite to millerite while leaving the copper sulfides intact, and to promote the reaction [28]: NiS + CuS + Cu S  + H S0  2  2  + J0  4  -> NiS0  2  4  + 3 CuS + H 0 2  (2-33)  autoclave leaching is carried out with only a slight oxygen overpressure. Depending on the ORP and pH, metallic copper reacts to form either cuprite or antlerite: 2 Cu + j 0  -> Cu 0  2  3 Cu + H 0  + H S0  2  2  + j 0  4  (2-34)  2  2  -> CuS0 -2Cu(OH) 4  (2-35)  2  The only sources of soluble copper are cuprite: Cu 0 2  + 2 H S0 2  4  + j 0  2  -> 2 CuS0  4  + 2 H0  (2-36)  2  and chalcocite, which may react in several steps to covellite under the mildly oxidizing conditions in the autoclaves: Cu S 2  + xH S0 2  4  + f 0  2  -> xCuS0  + Cu _ S  4  2  + H0  x  2  (2-37)  In the Outokumpu process lead is scavenged from solution through the addition of Ba(OH) , 2  which isomorphously co-precipitates P b S 0 with BaS0 . Cobalt and a host of other impurity 4  4  elements are removed from the pregnant N i S 0 solution by "group" precipitation with a slurry of 4  the powerful oxidizing agent "nickelic" hydroxide (Ni0 _ -nH 0) at near neutral pH and elevated 3  4  2  temperature. No detailed information is available yet about the separate iron removal step added recently to the Harjavalta flowsheet [30]. Nickelic hydroxide is produced electrolytically from a N i S 0  4  catholyte side stream.  Taking the formula of nickelic hydroxide as Ni(OH) for simplicity reasons, the reactions are: 3  NiS0  4  + 2 NaOH  -> Ni(OH)  2  +  Na S0 2  4  (2-38)  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice Ni(OH)  + H0  2  CoS0  4  ->• Ni(OH)  2  +H  +  3  + Ni(OH)  3  -> Co(OH)  3  24  + e  (2-39)  + NiSO.4  (2-40)  The cobalt cake is routed to the Kokkola Works, where cobalt and nickel powders are produced according to Sherritt's ammoniacal pressure leaching and hydrogen reduction technology. Nickel is electrodeposited from the all-sulfate electrolyte in divided cells with lead anodes. sodium and sulfuric acid are bled from the  anolyte  as  crystalline  Excess  "Glauber's salt"  (Na SO -10H O). The concentration of N a S 0 is kept at 150 g/L. 2  4  2  2  4  After start-up of the second autoclave, the expensive electrolytic removal of copper (as a powder) between the atmospheric leaching stages was abandoned. Instead, the copper sulfide-iron oxide autoclave residue can now be fed directly to the copper smelter. The P(G)M values of the copper anode slimes are ultimately refined at the Pori Works. 2.3.3.2 Zimbabwean operations Outokumpu has become a process licenser with a record of successful technology transfer. Variants of the Harjavalta process are practiced by the Empress Refinery and Bindura Refinery in Zimbabwe [31]. A third Zimbabwean refinery, the BHP-Delta Gold Hartley Platinum Complex [30], is scheduled to reach full production by early 1997.  2.3.4  Sherritt technology  2.3.4.1  Introduction  The original Sherritt autoclave process was developed in the early 1950's. It was designed to process the nickel-cobalt sulfide concentrates produced from an ore deposit at Lynn Lake, Manitoba, in an ammoniacal ammonium sulfate (AAS) medium.  The process, invented by  Professor Forward of U B C , and pioneered and perfected at the company's refinery in Fort Saskatchewan, Alberta, is well documented in literature and therefore not reviewed here. Ammonia has been employed extensively in the metallurgical industry, because (1) it is a strong and selective ligand for nickel, cobalt and copper, (2) it is readily recovered, and (3) it is compatible with carbon steel equipment.  In contrast, the use of pressure metallurgy has been  extended to acidic sulfate lixiviants only a few decades ago.  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  25  Today there are six matte leaching plants alone which utilize Sherritt's aqueous pressure oxidation technology.  The company's own facility now operates as a toll refinery treating a  variety of feed materials, such as matte, concentrates and precipitates, in an A A S medium to produce nickel and cobalt powder and briquettes, and ammonium sulfate. Although designed for a nickel concentrate feed, the Western Mining Kwinana Refinery in Australia [32] currently treats matte from the Kambalda smelter using a similar process.  The Impala, Rustenburg,  Western and Northam operations, all located within the Bushveld Igneous Complex in South Africa, employ acidic pressure leaching processes to separate the primary product, a P(G)M concentrate, from by-products nickel, copper and cobalt. Another South African acidic pressure leaching plant, the small-scale nickel-copper refinery of Barplats Platinum, was in operation only from 1989 to 1991. The complexity of the South African installations differs greatly with the production scale and the relative value of P(G)M's and base metals in the matte. In this section the flowsheets of the Impala and Rustenburg base metal refineries are described in detail. The smaller Western Platinum, Barplats Platinum and Northam Platinum operations are discussed briefly with the aid of a general, conceptual flowsheet.  2.3.4.2 Impala Platinum Acid pressure oxidation of matte was first put into production by Impala Platinum in 1969 at its nickel-copper refinery in Springs, Transvaal. Figure 2-7 is a simplified flowsheet of the plant. Ore from the Merensky Reef is mined, concentrated and smelted into a matte containing about 50% nickel, 28% copper, 0.5-1% iron and minor quantities of cobalt and P(G)M's. The main matte minerals are cobaltiferous heazlewoodite (Ni S ), nickeliferous chalcocite (Cu S) and 3  2  2  djurleite (Cu S ) , an alloy phase, and "refractory" oxide mineralization, including the spinels (9 3  chromite (FeCr 0 ), trevorite (NiFe 0 ) and magnetite (FeFe 0 ), as well as bunsenite (NiO) 2  4  2  4  2  4  dendrites in N i S [33]. The matte is granulated, wet-ground in ball mills, filtered, repulped with 3  2  spent electrolyte from the copper tankhouse and fed to a three-stage pressure leaching circuit.  The major portion of the nickel values (80-85%) is preferentially extracted at 280°F and 135 psig in the four-compartment first-stage autoclave . Initially, leaching is carried out under oxidizing conditions in the presence of air or 0 . Later, the admission of oxygen is discontinued to enable 2  dissolved copper to exchange with unleached nickel sulfide. The total retention time in the firststage autoclave is 3 hours.  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  _  H2SQ4  Ni powder  2  2  3rd stage leach  O2  T  jarosite residue  I Ni  Ni reduction  I  p  q+  Co powder  X  G  M  r e s i d u e  Cu EW  spent  Cu cathodes  evaporation lime^  ( ) Fe residue  I pregnantl Fe removal  Nhh  powder  Co reduction H2-  residue^ sol'n recycle)  H S04_  I  H2-  2nd stage leach  O2*  jarosite process  (NH4)2S04  Ithickener underflow  H S04^  sol'n adjustment NH3  _  cementate  Cu removal  O2  L  1 st stage leach  07  thickener overflow  _  26  lime boil gypsum tailings  Figure 2-7.  Simplified flowsheet of the Impala Platinum nickel-copper refinery (after [34,35]).  The alloy phase and cementate react rapidly in the feed tanks to the first-stage leach: M + H S0 2  4  + j 0  2  -» MS0  +  4  (2-41)  H0 2  In essence, first-stage leaching can be described by the following reaction chemistry [33,34]: - oxidative leaching: Ni S 3  2  2 Ni S 3  3 Ni S 3  2  2  + H S0  4  + j 0  2  -> NiS0  + 3 H S0  4  + j 0  2  -> 3 NiS0  2  + 2 CuS0  2  4  + 4 H S0 2  4  + 2 NiS +  4  + Ni S  + 2 0  4  2  -> 6 NiS0  3  4  + Ni S 3  2  + 3  4  4  (2-28)  H0  (2-42)  H0 2  + 2 CuS + 4 H 0 2  (2-43)  Chapter 2  27  Industrial Nickel-Copper Matte Leaching Practice 5 Cu S 2  Cu S 9  5  + j 0  + H S0 2  4  + 4 H S0 2  -> CuS0  2  + 2 0  4  + Cu S  4  -+'4 CuS0  9  2  4  5  +  (2-44)  H0 2  + 5 CuS + 4  H0  (2-45)  2  - non-oxidative leaching: NiS + CuS0  4  -> NiS0  4  (2-46)  + CuS  For simplicity of stoichiometry, the formula of digenite, Cu, S , is represented as C u S in the 7 6  9  5  above reactions. Leach liquor overflow from the first-stage leach thickener is subjected to two impurity removal processes.  First, copper is cemented from solution at pH 2-3 using nickel powder.  The  cementate is recycled to the first-stage leach. Secondly, the copper-free solution, now containing 100 g/L nickel, is treated in an iron removal stage, which has long been the major bottleneck in the Impala process [35]. Originally, the pH was adjusted with ammonia to 5.5 under aeration to precipitate iron as ferric hydroxide and arsenic as ferric arsenate.  The iron-arsenic residue so obtained,  unfortunately, also contained copious quantities of co-precipitated nickel and cobalt, and was returned to the smelter to reject iron and recover the sought-after base metals. Arsenic, due to accumulation in the integrated metallurgical flowsheet, reached intolerable levels in the nickel and cobalt recovery circuits. Hence, a separate leach circuit for the residue was developed, which involved a sequence of leaching in sulfuric acid of increasing strength and filtration for recovery of nickel and cobalt, and "bleeding o f f of iron and arsenic in a goethite residue. The great number of disadvantages inherent to this leaching process finally led to the implementation of a jarosite process by the late 1980's.  Most of the iron, arsenic and lead now leave the  hydrometallurgical circuit in a mixed alkali jarosite-ferric arsenate residue.  Filtered solution from the jarosite precipitation stage is prepared for hydrogen reduction by the addition of recycled ( N H ) S 0 and N H . Nickel powder is precipitated from the resulting 4  2  4  3  ammoniacal nickel "diammine" sulfate solution and either sold as such or briquetted. Cobalt was previously sold in the form of a mixed nickel-cobalt sulfide obtained from the nickel reduction end solution with NaHS, but since 1982 is recovered separately using an altered version of the cobaltic aquopentammine route. Regeneration of the barren A A S stream regenerates ( N H ) S 0 (evaporator) and N H 4  (lime boil) for recycle to the solution adjustment step:  2  4  3  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice Ca(OH)  + (NH ) S0  2  4  2  ->• CaS0 -2H 0  4  4  28  + 2 NH  2  (2-47)  3  The gypsum formed in reaction (2-47) is pumped to a tailings pond. The first-stage leach residue is subjected to a second T O L to extract as much of the nickel, copper, cobalt and iron values as possible.  The leaching conditions are 275°F, 20 psig partial  pressure of oxygen and 4 hours residence time. The main second-stage leaching reactions are (2-45) and: MS + 0  -»  2  Ni S 3  4  + H0 2  + If 0  2  (2-48)  MS0  4  -> 3 NiS0  4  +  (2-49)  H S0 2  4  with M = Cu or Ni. Prior to becoming the feed to the copper tankhouse, iron is removed from the second-stage filtrate by oxidation-hydrolysis at pH 2.8 and 180°F.  Due to the presence of  some co-precipitated copper and nickel, the iron residue is returned to the smelter. Spent copper electrolyte is recycled to the first-stage leach. The second-stage residue is further treated in a similar third acidic oxidative leaching process to provide a final residue rich in P(G)M's. This residue constitutes the feed to the Impala P G M Refinery. The third stage leach liquor is returned to the second-stage autoclave.  2.3.4.3 Rustenburg Base Metal Refiners The Rustenburg Base Metal Refiners (RBMR) plant [36,37,38] consists of a P G M enrichment facility and a base metal refinery serving various mines, concentrators and smelters in the Rustenburg district of Transvaal.  Hydrometallurgical operations at the integrated Rustenburg  Platinum Mines complex, the world's largest producer of PGM's, commenced in 1966.  A new  refinery with a Sherritt designed leaching circuit was built and commissioned by the end of 1981.  Converter matte, containing about 43% nickel, 29% copper, 23% sulfur, 1.5% iron and 0.5% cobalt, is cast into ingots at an adjacent smelter and slow-cooled for a period of three days. Following crushing and fine milling, matte is transported to R B M R where it is subjected to magnetic separation. The magnetic alloy fraction then undergoes hydrometallurgical treatment, in order to produce an enriched P(G)M concentrate as a feedstock for the Precious Metals Refinery. The resulting leach solution is pumped to the primary pressure leach, whereas the nonmagnetic sulfide fraction of the matte is advanced to the atmospheric leaching stage.  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  HP non-magnetic Ni-Cu malt*  MATTE aala  Cu tor  Cu bu l l i o n to a m s l t e r  ntagnatic c o n e , laachate  air  29  solution  COPPER REMOVAL  neutral spent electrolyte  PRIMARY LEACH  PYROMET PLANT  SECONDARY LEACH  Hue sea  copper scrap  LEAD REMOVAL  barium hydrox Ida "  aulphurio formalin Nl/Co caka  COBALT REMOVAL  CAKE TREATMENT  FINAL FILTRATION NICKELIC CELLS  Ni tor  nlckal anotyta  acid I  SO SCRUBBER  aulphita  solution  CAKE DISS. AND PURIF.  FINAL FILTRATION  SOLVENT EXTRACTION  COPPER E-WINNINQ  CATHODES aala  Se cake  cauatic aoda  Cu CATHODES tor ami*  NICKEL E-WINNINQ  caualle aoda  SELENIUM REMOVAL  ,, Pb caks to amoltori  CRYSTAL. PLANT  COBALT for aala  SULPHATE  CRYSTAL. PLANT  SODIUM tor aala  SULPHATE  nickal anolyta aolvent e x t r a c t i o n raftlnata  NI HYDROXIDE DISSOLUTION  NI HYDROXIDE PRECIPITATION cauatic aoda  Figure 2-8.  Simplified flowsheet of the Rustenburg Base Metal Refiners (RBMR) process [36].  Figure 2-8 is a simplified flowsheet of the R B M R process [36,37]. Fresh matte is contacted with a mixture of primary pressure leach discharge solution and "sulfur removal" solution in a series of four aerated continuous stirred tank reactors. The reactors have a volume of 170 m each and 3  the total residence time is 12 hours.  The atmospheric copper removal process produces a  virtually acid-, copper- and iron-free solution, while stoichiometrically equivalent amounts of nickel pass into solution. Contrary to the Outokumpu process, copper is precipitated by the alloy-free matte either via metathesis with N i S at a solution pH below 2.5: 3  Ni S  2  3  2  + 2 CuS0  4  -> 2 NiS0  4  + M S + Cu S 2  (2-50)  or through hydrolysis at a solution pH higher than 2.5. Excess acid is simultaneously removed by the oxidation of heazlewoodite (reaction (2-28)), which really proceeds stepwise through the initial formation of godlevskite, N i S : ?  6  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice 3 Ni S 3  + 2H S0  2  2  3 Ni S 7  + 0  4  2  4  + Ni S  4  + j 0  + H S0  6  -> 2 NiS0  2  7  -> NiS0  2  30  + 2 H0  6  2  + 6 NiS + H 0  4  2  (2-51) (2-52)  Reactions (2-51) and (2-52) lower the solution acidity, causing precipitation of antlerite in the pH range 2.5-4.5: 3CuS0  + 4H 0  4  -> CuS0 -2Cu(OH)  2  4  2  + 2 H S0 2  4  (2-53)  and of ferric hydroxide or parabutlerite in the range 4.5-6: Fe (S0 )  + 6 H0  -> 2 Fe(OH)  Fe (S0 )  + 2H 0  -> 2 FeS0 (OH)  2  2  4  4  3  3  2  2  3  4  + 3 H S0 2  4  + H S0 2  4  (2-54) (2-55)  The acid released in the above hydrolysis reactions is consumed by unreacted heazlewoodite and godlevskite. Good aeration is required to promote hydrogen ion removal by reaction (2-51) and (2-52) and shift the equilibrium of reaction (2-53) in favor of antlerite precipitation. Hydrolysis is completed at pH 6, when the atmospheric leaching discharge solution contains less than 10 mg/L each of copper and iron. Iron rejection is the rate limiting process of the atmospheric leach. Hence, the operating temperature is chosen at 75-80°C where the iron removal rate reaches an optimum. The presence of cupric ion in the feed solution is essential for efficient iron removal. Although the kinetics of copper rejection through metathesis (reaction (2-50)) are much faster, this mode of operation seems to be unsuitable for mattes with a high iron content. The reason for this is an unacceptably long iron hydrolysis in such instances.  Therefore, R B M R  operate the copper removal stage at low acidity, necessitating preneutralization of the primary leach liquor to pH 2.5-3 with caustic soda. In this way the kinetics of copper hydrolysis alone determine the outcome of the whole operation, that is, final pH, copper and iron concentrations.  The atmospheric leach slurry is discharged into a large thickener with buffering capacity in the case of process upsets. The impure N i S 0 overflow solution is purified in three steps prior to 4  nickel electrowinning.  First, lead is removed with Ba(OH) .  using the Outokumpu nickelic hydroxide process. electrolyte then enters the tankhouse.  2  Secondly, cobalt is precipitated  After fine filtration, pregnant nickel  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  31  While the lead cake is recycled to the smelter, the cobalt cake is further treated for cobalt recovery.  It is first leached with spent nickel electrolyte to redissolve co-precipitated nickel  hydroxides. After filtration the enriched cake is then subjected to a reducing leach in a sulfuric acid-formaldehyde solution. Following solution purification, cobalt is extracted using D2EHPA. Finally, the cobalt sulfate strip solution is crystallized to C o S 0 - 7 H 0 for sale. 4  The nickel-  2  containing raffinate is mixed with spent nickel electrolyte and pumped to the "sulfur removal" stage. In the "sulfur removal" stage the nickel stream is split into two stoichiometric proportions. One part is neutralized to pH 9 with NaOH. The precipitated Ni(OH) is filtered on 2  drum filters and redissolved in the remaining volume of spent electrolyte. Since the redissolved cake contains some lead (from anodes) and copper (from busbars) generated during nickel electrowinning, it is recycled to the copper removal stage. The neutralization filtrate is processed in the N a S 0 plant into salable anhydrous crystals. 2  4  Thickener underflow slurry is pumped to the primary pressure leach at 135°C and 1050 kPa in a four-compartment horizontal Sherritt autoclave with a working volume of 80 m . Spent copper 3  electrolyte is used as the source of acid. Only the leading two compartments are aerated to ensure maximum nickel solubilization; the last two compartments are operated without air to reprecipitate part of the copper in solution.  The total residence time in the primary leach  amounts to 4 hours. Following liquid-solid separation, the primary leach residue is advanced to the secondary leach.  About 80-85% nickel extraction is achieved in the combined copper  removal and primary leach stages. After feed preparation and preheating, the atmospheric solids are essentially a mixture of djurleite (Cuj S ) , taken as Cu S for simplicity reasons, and millerite. 93  2  The major reactions  taking place during the oxidizing period are dissolution of nickel from millerite: + j 0  4 NiS + H S0 2  4  NiS0  2  + Ni S  4  3  4  + H0 2  (2-29)  and copper from djurleite: Cu S 2  + H S0 2  4  + \ 0  -> CuS0  2  + CuS + H 0  4  2  (2-56)  Furthermore, a substantial part of the nickel passes into solution by direct oxidation of millerite: NiS + 2 0  2  -»  NiS0  4  (2-57)  Chapter 2  Industrial Nickel-Copper  Matte Leaching Practice  32  During the non-oxidizing period a large quantity of copper is precipitated through metathesis: NiS + CuS0  -> NiS0  4  + CuS  4  (2-46)  The drop in pH observed in this stage is believed to be due to a reaction between polydymite and cupric ion. Although the actual mechanism likely involves the formation of a range of nonstoichiometric metal sulfides prior to the formation of a discrete digenite (Cu S ) phase, the 9  overall reaction can be conveniently represented by: 4Ni S 3  4  + 9CuS0  + 8H 0  4  -» 3 NiS0  2  4  5  • + 9 NiS + Cu S 9  5  + 8 H S0 2  4  (2-58)  In addition, digenite may be formed in-situ by a solid phase interaction between the covellite surface layer and the residual djurleite core: 4 Cu S + CuS -> Cu S 2  9  (2-59)  5  Clearly, it is important the nickel extraction be limited during the oxidizing period to leave a sufficient quantity of nickel sulfides for copper precipitation. Secondary pressure leaching is carried out at 145°C and 1050 kPa in autoclaves similar in construction to the primary ones. Again, spent copper electrolyte is used as the source of acid. Aeration with oxygen-enriched air [38] ensures complete dissolution of copper (98%) and nickel (99%). The overall retention time is normally 8 hours. The major reaction in the initial phases of oxidation is the decomposition of digenite: Cu S 9  5  + 4H S0 2  + 20  4  -> 4 CuS0  2  + 5 CuS + 4 H 0  4  2  (2-45)  The generated covellite is either directly oxidized to copper sulfate: CuS + 2 0  2  -> CuS0  (2-22)  4  or can be decomposed to elemental sulfur:  CuS + H S0 2  4  + j 0  2  -> CuS0  4  + S° +  H0 2  (2-60)  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  33  Reactions (2-45) and (2-22) occur simultaneously, oxidized covellite being substituted by covellite generated from digenite, so initially no change occurs in the CuS content of the solids. As the dissolution of copper subsulfides (reaction (2-36)) is kinetically very rapid, the overall leaching time is determined by the oxidation of covellite (reaction (2-22)). The decomposition of CuS according to reaction (2-22) is catalyzed by iron. Residual polydymite dissolves according to the overall stoichiometry: Ni S 3  4  + H0 2  + y- 0  2  -> 3 NiS0  4  + H S0 2  (2-49)  4  However, the reaction sequence for the decomposition of (copper-bearing) polydymite remains obscure.  Millerite is partially converted to polydymite (reaction (2-29)) and partially directly  oxidized to N i S 0 (reaction (2-57)). 4  Iron is precipitated in the form of hematite and sodium jarosite (NaFe (S0 ) (OH) ). 3  4  2  6  After filtration and washing, the final iron-rich leach residue, the only outlet for iron entering the R B M R process, is treated pyrometallurgically to separate it into a copper bullion containing residual P(G)M's and gold, and a copper matte with silver and other, mostly non-metallic impurities. The bullion is recycled to the convertors, whereas the matte is sold to an overseas refiner. The copper-rich filtrate is diluted with spent electrolyte to avoid crystallization of C u S 0  4  and treated with S 0 in Sherritt's patented selenium removal process. Selenium cake, due to its 2  P G M content, is combined with the secondary leach residue for further treatment. No further purification other than solution clarification is required in the copper circuit, and the solution, now containing 75 g/L copper, 30 g/L nickel and 60 g/L acid, enters the copper tankhouse. As mentioned, spent copper electrolyte is recycled to both pressure leach stages to supply acid for the dissolution of the base metals.  2.3.4.4 Western, Barplats and Northam Platinum The Sherritt nickel-copper matte acid leach process in its most recent state of development employs two-stage rather than three-stage leaching. Furthermore, the use of autoclaves has been minimized to just one stage. Figure 2-9 is a conceptual flowsheet of the process practiced at the Western Platinum [39] and Northam Platinum [40] base metal refineries. Barplats Platinum only shortly operated a similar process [40,41].  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  34  Ground Matte  H2O H2SO4  I  0  2  PGM Residue  Se Residue  Cu Cathode  Figure 2-9.  Conceptual flowsheet of the Western, Barplats and Northam Platinum nickel-copper refineries [41].  Although a first-stage pressure leach provides an improved separation of nickel from copper and thus permits the production of higher grade copper cathode, atmospheric leaching at 85-90°C in a cascade of agitated tanks is preferred for the small-scale operations discussed in this section. In the interests of simplicity and low capital costs, relatively pure N i S 0 - 6 H 0 for upgrading 4  2  elsewhere is crystallized from the atmospheric leach solution. The second-stage leach, which is designed to yield maximum extraction of nickel, copper and sulfur while leaving the P(G)M's intact, is a pressure leach operating at 150-160°C under 150-350 kPa partial oxygen pressure.  After pressure filtration, the second-stage leachate is  treated with S 0 to eliminate selenium prior to copper electrowinning. Due to the higher nickel 2  content of the copper electrolyte, copper electrodeposition is typically carried out at a somewhat lower current density. Spent copper electrolyte is recycled to provide the acid required in the two leaching stages.  The P(G)M's are recovered as a high-grade concentrate in the second-stage  residue and in the copper selenide precipitate emanating from the selenium removal process.  Chapter 2 2.3.5  35  Industrial Nickel-Copper Matte Leaching Practice  Mansfeld Kombinat technology  All of the previous sections deal with the hydrometallurgical processing of primary nickel-copper matte. The Nickelhiitte Aue of the V E B Mansfeld Kombinat in former East Germany operates a leaching process for the treatment of a secondary matte produced from several secondary and internally generated revert materials. The final products are cathode nickel or N i S 0 , cathode 4  copper and cobalt oxide.  The following paragraph compiles the limited information available  about the leaching practice [42,43]. The matte is ground to 100% minus 100 um, and subjected to batch pressure leaching at 130°C and 0.5-1 MPa partial oxygen pressure in one of three 20 m autoclaves. The working volume of 3  each autoclave is 10 m , and the total duration of the heating-leaching-cooling cycle is 12 hours. 3  Oxygen is introduced through the hollow shaft of the gas pumping impeller. The leach filtrate typically contains 50-80 g/L nickel, < 15 g/L copper, < 3 g/L cobalt, 2-5 g/L iron and 20-30 g/L H S 0 . To avoid pressure filtration problems, low matte iron contents are preferred. 2  2.4  4  Nitrate Lixiviants  Matte leaching in nitrate lixiviants has never been commercialized, however, it was investigated extensively by S L N in France in the mid 1970's. Before the company finally adopted the allchloride route described in section 2.2.2, various hydrometallurgical processes for nickel production were evaluated, including two process variants which utilized nitric acid in the leaching stage [44]. The common first step of both piloted processes comprised the oxidative leaching of matte from New Caledonia, containing 75% nickel, 0.04% copper, 20% sulfur, 2.5% iron and 1.9% cobalt, in a nitric acid solution at atmospheric pressure and 90°C.  Nitrous gases generated during  leaching were reconverted to FIN0 . The principal leaching reactions were: 3  Ni + 2HN0  3  Ni S 3  2  + 2HN0  3  + H  - » 2 NiS0  + Ni(N0 )  3  + 40  2  (2-61)  -> Ni(N0 )  2  4  2  3  2  +  H  2  (2-62)  Since nitrate had to be removed from the leach solution prior to subsequent process steps, leaching was followed by a sulfation step. Here, sulfuric acid was added to the slurry in order to displace nitrate ion and satisfy the stoichiometry:  Chapter 2  Industrial Nickel-Copper Matte Leaching Practice  uNi S"  + H S0  2  2  + 20  4  -> 2 NiS0  2  4  36  + H  (2-63)  2  wherein "Ni S" represents the average overall composition of the matte. 2  The crude nickel sulfate-nitrate solution was purified by simultaneous hydrolysis of iron and precipitation of copper and zinc with H S at pH 4. Cobalt was precipitated using either nickelic 2  hydroxide or NaClO. The pure N i S 0 - N i ( N 0 ) solution was then subjected to a denitrification 4  step. NiS0  3  2  Nitrate separation was carried out either "directly" through selective crystallization of 4  or "indirectly" by precipitation of a basic nickel carbonate (BNC).  Electrolyte was  prepared from the nitrate-free nickel feed by redissolution of the nickel sulfate crystals in water or by the dissolution of B N C in sulfuric acid: NiC0  3  • 3Ni(OH)  2  + 4H S0 2  4  -» 4NiS0  4  + C0  2  +  7H 0 2  (2-64)  Finally, nickel was recovered through electrowinning. Both process variants were abandoned because of high operating costs due to severe operating difficulties, including high nickel losses in the purification stage, insufficient cathode impurity and unacceptably high anode corrosion rates.  Chapter 3  37  Development of Pourbaix Diagrams CHAPTER 3  DEVELOPMENT OF POURBAIX DIAGRAMS FOR THE SYSTEMS ARSENIC-SULFUR-WATER A N D NICKEL-SULFUR-WATER  3.1  Introduction  Pourbaix diagrams, despite their limitations, are excellent tools for the  study of the  thermochemistry of metal-ligand-water systems. In this chapter, therefore, an attempt is made to develop potential-pH diagrams for the system arsenic-sulfur-water at 25 and 160°C, the latter being the temperature at which the First-Stage Leach (FSL) process at INCO's Copper Refinery currently operates.  For reasons that will become clear in Chapter 5, an E - p H diagram for the h  system nickel-sulfur-water at room temperature is also computed. Species sets, thermodynamic data and the framework of the the spreadsheet calculations pertaining to the high-temperature arsenic diagram are included in Appendix A.  The diagrams are generated with the aid of the powerful CSIRO Thermochemistry System (Version V) software. For aqueous species, the recent IUPAC thermochemical data compiled by Bard et al. [45] are preferred over the older NBS values by Wagman et al. [46]. Additional free energy data for aqueous arsenic and nickel species are taken from Robins [47] and Pourbaix [48], respectively.  Heat capacity data for hydrogen, oxygen, arsenic and sulfur - in the form of  o polynomial equations relating C to temperature - are taken from Rao [49]. Barin's volumes [50] p  are utilized as a source of standard entropy values of pure compounds at 25°C.  Likewise, the  Gibbs free energies of formation of all mineral phases, gases and water are obtained from Barin's tables either directly or by linear interpolation. 3.2  Sulfur-Water Diagrams  Although a complete review of the theory of Pourbaix diagrams is beyond the scope of this thesis, the system sulfur-water warrants special attention. For thermodynamic analysis of many hydrometallurgical systems involving sulfur as a ligand, the great number of aqueous sulfur species can conveniently be reduced to hydrogen sulfide and its ionization products, rhombic sulfur, and bisulfate and sulfate ions. Figure 3-1 is the "standard" S - H 0 diagram at 25°C and unit activity of those species. The host of - mostly metastable - ions 2  Chapter 3  Development of Pourbaix Diagrams  38  not considered here includes thiosulfate, polythionates, polysulfides, peroxydisulfate as well as dissociation products of sulfurous and "Caro's acid". Z. Vi  PLOT L A B E L S T = 298.15K IS] = 1 h 5THBLE RRER5 9 B C  H2 S IG) H 5 <-> (RQ) S <2-> IRQ]  D E F  5 H S 04 <-> (RQ) S 0 4 <2-> IRQ)  H20 S T A B I L I T Y 1 OXYGEN 2 HTDRQCEN  -2  0  2  4  6  8  10  L2  14=  LIMITS  IB  P H  Figure 3-1.  "Standard" sulfur-water diagram at 25°C with 1 atm gas pressure and unit activity of all solutes.  Although realistic in the field of geochemistry when the lines in the diagram are likely to present true equilibria reached after millions of years, Figure 3-1 is largely unuseful in hydrometallurgy. It is well known that elemental sulfur is not oxidized to sulfate by ferric ion at ambient conditions.  Relying on such experimental hindsights, the "stability" (really metastability) of  elemental sulfur may be artificially increased by 300 kJ/mole (after E. Peters). Since the sulfursulfide equilibria must not be disturbed, this is done by making the Gibbs free energies of formation of the sulfate and bisulfate ion less negative by this amount.  The lines for the  "equilibria": S0  2  + 8H  + 6e  +  4  <-> 5 ° + 4 H 0  (3-1)  2  and: HS0  + 7H  +  4  + 6e  <-> S° + 4  H0 2  (3-2)  Chapter 3  39  Development of Pourbaix Diagrams  are thus shifted upwards in the diagram by 0.5 V (50 kJ/eq). The result of this operation can clearly be seen in the "extended" sulfur-water diagram of Figure 3-2. 2.  1  1  1  \  1  1  -. ;:  LRBELS  T = 298.15K IS]  1. 5  E  5TRBLE F  1.  H2 S (G)  C  S <2-> IRQ)  D E F  ~" — _  1.5 — """"---^  RRER5  R B  c JJ  = I M  H S <-> (HQ]  5  H S 04 <•> IRQ) S CM <2-> IRQ)  H20 S T A B I L I T Y  0  LIMITS  1 OXYGEN 2 HYDROGEN  -3n B  •1.1 1  1. 5  2  i  1  i  1  C  1  114  12  16  P H  Figure 3-2.  "Extended" sulfur-water diagram at 25°C with 1 atm gas pressure and unit activity of all solutes.  In accordance with the practice in established works, particularly Pourbaix's Atlas [48], the pH scale in the sulfur-water diagrams ranges from -2 to 16. It is noted, however, that these lower and upper pH values represent physically impossible acid and alkali concentrations, respectively, and are therefore really thermodynamic fiction.  Moreover, at very high acid or alkali  concentrations the concept of the "aqueous" system no longer holds.  3.3  Arsenic-Sulfur-Water Diagrams  Arsenic, like sulfur, forms a large number of species in aqueous solution.  Considered in the  present analysis of the A s - S - H 0 system are arsine gas (AsH ), the As(III) arsenyl ion (AsO ), as +  2  3  well as meta-arsenious acid (HAs0 ), ortho-arsenious acid ( H A s 0 ) and arsenic acid ( H A s 0 ) 2  3  3  3  4  and their deprotonation products. Left out of the calculations are other arsenic hydrides, such as diarsane (As H ), the As(V) arsenyl ion ( A s 0 ) , the perarsenyl ion ( A s 0 ) and all thioarsenite 2  4  and thioarsenate species.  2  3  Chapter 3  40  Development of Pourbaix Diagrams  The only solid phases included in the development of the diagrams are the grey, rhombohedral allotropic form of metallic arsenic, claudetite (monoclinic A s 0 ) , realgar (AsS) and orpiment 2  3  (As S ). 2  3.3.1  3  Ambient temperature  Figure 3-3 shows the extended A s - S - H 0 diagram at 25°C with 1 atm gas pressure and unit 2  activity of all solutes. PLOT L A B E L S T = 298.15K [Rsl  = 1 M  IS] = 1 M  STRBLE RRERS A  As H3 IC)  B  Ac  D  A s 2 S3  C E F  C  H I  J  K  As 5  A s 0 3 <3-> (RQl A3 OH <3-> (RQ)  A s 0 <+>  (AQ)  H3 A s 014 (AQ)  H2 Aa 04 <•> (RQ]  H A s 0M <2-> (AQ] As2 03  H20 S T A B I L I T Y 1 OXYGEN  LIMITS  2 HYDROGEN  P H  Figure 3-3.  Extended arsenic-sulfur-water diagram at 25°C with 1 atm gas pressure and 1 molal activity of aqueous arsenic and sulfur species.  The number of unstable aqueous species at the conditions for which Figure 3-3 is generated is limited to monomeric meta-arsenious acid and its dissociation product, and the mono- and diprotonated ortho-arsenious acid ions. Realgar does not appear in the diagram when Barin's Gibbs free energy of -133.004 kJ per mole A s S is used. Since the mineral is commonly found in natural deposits, this value seems 4  4  to be unreliable. Indeed, mineralogical thermodynamic data compiled by Robie et al. [51] and Vaughan and Craig [52] list a free energy of approximately -71 kJ per mole AsS. The adopted USGS value of -70.32 kJ/mole does produce the expected rim of realgar below the A s S region. 2  3  Chapter 3  41  Development of Pourbaix Diagrams  Unfortunately, no high-temperature USGS data are available, and a Barin value will be used in the A s - S - H 0 diagram at 160°C. 2  Upon lowering of the metalloid activity to 10" molal, the stability region of A s 0  is  3  2  "replaced" by that of H A s 0 . 3  3  3  Moreover, a window for the di-ortho arsenite ion then appears,  which, like the areas of predominance of all other aqueous arsenic species, grows at the expense of the orpiment domain when the arsenic solute activity is further decreased.  3.3.2  Elevated temperature  In order to calculate the high-temperature A s - S - H 0 diagram, the Gibbs free energies of 2  formation of all aqueous species at 160°C have to be computed using the Criss and Cobble theory [53,54,55]. The Criss and Cobble theory was developed in the mid 1960's.  It is based on the so-called  entropy correspondence principle. This principle implies that a standard state can be chosen at every temperature such that the entropies of a certain category of ions at that temperature are linearly related to their corresponding standard entropies at some reference temperature (usually 25°C).  The standard state has to be chosen properly by fixing the entropy of H  temperature [53].  +  at each  Once the entropy of an ion at elevated temperature has been estimated, the  average value of its heat capacity can be calculated [54]. The change in heat capacity of the ion o  formation reaction is then combined with values of AS SK 29  calculate  o  a n Q l  ^f,298K  f°  reaction to  r t n a t  o  AGj- . TK  Two problems arise when the empirical theory is applied to the present system: (1) unknown standard ionic entropies, and (2) classification of oxy-cations, complex non-oxy-anions and undissociated species. The Criss and Cobble calculations require an initial value of the standard entropy of each aqueous species.  Unfortunately, for the arsenyl, di-ortho-arsenite and ortho-arsenite ions that  number is not available and needs to be estimated. A value of 12.52 J/mole/K for S°  (AsO ) +  298K  is estimated using Latimer's method [56].  The standard entropies of the mono-protonated ortho-arsenite and ortho-arsenite ions are estimated on the basis of the dissociation series of arsenic acid. It is assumed that the stepwise decrease in entropy for the ortho-arsenious acid dissociation series is proportionally the same as 2  0  for arsenic acid. This engineering approach gives -35.1 and -248.9 J/mole/K for S and S°  (As0 ~), respectively, as rough starting values for the calculations. 3  298K  3  2  9  8  K  (HAsOf)  Chapter 3  42  Development of Pourbaix Diagrams  Within the framework of the Criss and Cobble theory, ions are classified as simple cations, simple anions (and OH"), oxy-anions and acid oxy-anions.  Therefore, likely erroneous  assumptions have to be made for cationic hydrolysis species such as the arsenyl ion, and acid anions like the bisulfide ion. Since there is some evidence that the entropies of oxy-cations such as the uranyl ion ( U 0 2 ) lie near the correspondence diagram for simple cations [53,55], A s O is assumed to be +  +  part of this category in the spreadsheet calculations. Bisulfide ion is classified as a simple anion in the present analysis. Undissociated species, such as the weak As(ni) and As(V) acids, have no charge and cancel Criss and Cobble's distinction between the "conventional" and "absolute" entropy scales. Neutral aqueous complexes, therefore, must be treated separately according to Helgeson's theory [57]. Like Criss and Cobble, Helgeson utilizes entropies as an entry into the problem of predicting thermochemical properties in solutions at elevated temperatures. In his theoretical interpretation of an isothermal entropy cycle for an arbitrary dissociation reaction, Helgeson separates entropies into hypothetical electrostatic and non-electrostatic terms. Thus, the temperature dependence of the thermodynamics of dissociation for complexes is described in terms of functions involving the dielectric constant of water and a power series consistent with non-electrostatic interaction in the absence of a dielectric medium. Although the concept of his theory may be abstract, Helgeson's equations are quite practical to use after the proper assumptions. Furthermore, they are fully compatible with the Criss and Cobble calculations. Figure 3-4 shows the extended A s - S - H 0 diagram at 160°C with 1 atm gas pressure and unit 2  activity of all solutes. The equations used in the development of this figure are summarized in Appendix A.  Many of the calculated free energy values at 160°C have been verified by  comparison with those computed at 150°C by Barner and Scheuerman [58].  In the interpretation of Figure 3-4 it should be borne in mind that the concept of the "neutral pH" condition changes with temperature due to variation of the dissociation constant of water. It can easily be calculated that pK  w  is in the vicinity of 11.5 at 160°C.  approximately 5.75 for equimolal activities of protons and hydroxyl ions.  This gives a p H of  Chapter 3  43  Development of Pourbaix Diagrams  2. 0  PLOT  LABELS  T = 433.15K IRs|  1.5h  IS] J  STABLE  1.0 0.5h 0  M  AREAS  R  Rs H3 (G)  B  Ae  C  R s 2 53  D  Rs 0 3 <3-> (HQ)  E  H fls 04 <2-> IRQ)  F  Ra  G  As 0 <+> (AQ)  H  -0.5  1  = 1 M =  CM <3>  (HO)  H3 Rs 0 3 ( f l u )  ]  H3 Rs OH (AQ1  J  H2 As  m <->  H20 S T A B I L I T Y  IRQ)  LIMITS  1 OXYGEN 2 HYDROGEN  1.0 1.5 "-2 1 -  0  2  14  6  8  10  12  14  16  P H  Figure 3-4.  Extended arsenic-sulfur-water diagram at 160°C with 1 atm gas pressure and 1 molal activity of aqueous arsenic and sulfur species.  The (sulfur) ligand equilibria must be interpreted with caution. Many uncertainties surround the ambiguous ionization of the bisulfate ion. Since the association of protons and sulfate weakens with increasing ionic strength, the bisulfate buffer point for practical solutions at ambient temperature is probably closer to pH 0 than 2 (after E. Peters). Comparing the stability regions of bisulfide ion in Figure 3-3 and 3-4, it shows that HS" becomes an increasingly strong acid as the temperature rises. At very high temperature it actually becomes a stronger acid than hydrogen sulfide and ceases to be an important dissolved sulfur species [55].  Realizing that the size and location of predominance areas of the arsenic species in the A s - S - H 0 2  diagram at elevated temperature are only as accurate as the thermodynamic data they are the result of, it appears that the arsenic oxide is no longer stable at the conditions of Figure 3-4. Besides the generally steeper lines in the diagram, the large stability regions of the arsenate ion and its di-protonated counterpart are also striking features of the high-temperature system. As mentioned before, the disappearence of the realgar band can be attributed to questionable thermodynamic data.  Chapter 3  44  Development of Pourbaix Diagrams  As for the status of high-temperature aqueous thermodynamics, the Criss and Cobble theory seems to be one of the few practical ones available. It has become evident in this chapter that Criss and Cobble's theory requires revision.  Despite its age and non-fundamental character,  however, it remains helpful in the study of high-temperature systems in hydrometallurgy. Greater accuracy in the calculation of thermodynamic data at elevated temperature, although desirable from an academic point of view, is likely trivial to industrial processes given the variations in the temperature at which reactors such as the FSL autoclave actually operate.  3.4  Nickel-Sulfur-Water Diagram  Figure 3-5 shows the extended N i - S - H 0 diagram at 25°C with 1 atm gas pressure and unit 2  activity of all solutes. It is generated with the reducing conditions after copper depletion of a FSL batch in mind. This confines the aqueous nickel species of interest to those formed by the metal in its divalent state. In addition to N i , N i O H (high acidity) and H N i 0 2 +  +  2  (high alkalinity)  are also considered. The solid phases included in the development of the diagram are bunsenite (NiO), millerite (NiS) and heazlewoodite (Ni S ). 3  2  PLOT LRBELS  1  T = 298.15K INM = M IS] = 1 M STRBLE RRERS H  Nt  C  Ni S  E  Ni 0  B  D  N i <2+> (HQ] N ; 3 S2  H20 S T A B I L I T Y 1 OXYGEN 2 HYDROGEN  LIMITS  PH  Figure 3-5.  Extended nickel-sulfur-water diagram at 25°C with 1 atm gas pressure and 1 molal activity of aqueous nickel and sulfur species.  The behavior of nickel at low potentials will be adressed in Chapter 5.  Chapter 4  45  Experimental Work CHAPTER 4  EXPERIMENTAL  4.1  WORK  Sampling Program  Since IPC residue is highly sensitive to oxidation by air, special measures have to be taken to preserve the freshness o f samples o f the carbonylation residue during long-term storage. O n the basis o f INCO's experience with various methods o f sample preparation and storage, it was decided to freeze the samples as wet cakes for the present investigations.  Figure 4-1.  Sample taking from the sample point o f the IPC thickener underflow line.  Chapter 4  Experimental Work  46  In May, 1994, a sampling campaign was undertaken at INCO's C R E D plant in Copper Cliff, Ontario, to obtain representative IPC feeds for both the preliminary test work on site and the leaching program to be carried out at U B C . During the 16-day period from May 9 to 24, at least four slurry samples of approximately 500 mL each were drawn daily from the sample point of the IPC thickener underflow line depicted in Figure 4-1. Sampling was done in the early morning. Two by two the bottle contents were quickly filtered over Buchner funnels under vacuum and washed thoroughly with distilled water. The IPC filter cakes were then transferred onto watch glasses. A head sample was composed of several spatula scoops from all cakes. Bits of material were selected from several parts of the cakes and, in view of possible particle segregation within the cakes, care was taken to extract the entire cake thickness. The filter cakes were put in nitrogen-purged double plastic bags, labeled and temporarily stored on ice in a cooler. Every 2 or 3 days all daily samples from the cooler were shipped to a freezer at the Clarabelle Mill. The head sample was repulped in distilled water, filtered under vacuum and washed with approximately 250 mL of methanol. To ensure effective displacement of residual water in the cake pores, methanol was added just before cake cracking commenced. The methanol-treated cake was transferred quantitatively onto a watch glass and dried in a vacuum desiccator. The dry solids were submitted for assay at INCO's Copper Cliff Central Process Technology Laboratory.  The samples were trucked to U B C in frozen state on May 25. They arrived in good shape in Vancouver on May 30 and were immediately stored in a freezer in the Corrosion Laboratory. A l l 66 samples were kept congealed until use in the leaching program.  4.2  Leaching Program  4.2.1  Scope  The F S L process parameters investigated in the present leaching program are summarized in Table 4-1.  Chapter 4  47  Experimental Work  Table 4-1.  Summary of First-Stage Leach (FSL) parameters investigated in the leaching program.  Physicomechanical parameters  agitation rate pulp density temperature retention time  Chemical parameters  composition of IPC residue  Cu : S  composition of electrolyte  copper concentration  t t  ratio  acid concentration effect of organic additives effect of arsenic  In addition to the 66 autoclave tests carried out as part of this program, a number of releaching experiments were performed with FSL residues of the copper-depleted source in two different media. The purpose of those tests was to acquire more specific information on the behavior of arsenic. Earlier proposed autoclave runs with IPC residues "doped" with particular mineral phases such as pentlandite ((Ni,Co,Fe) S ), trevorite (NiFe 0 ) and domeykite (Cu As) were canceled 9  8  2  4  3  when it became apparent that specimens of those minerals of acceptable purity were not readily available.  After reconsideration, plans for a series of experiments involving air discharge of  pressure leached slurry were also dropped.  4.2.2  Batch make-up  The "recipe" for the first-stage leaching tests was obtained through downscaling of the actual FSL process (Appendix B), taking into account information from shift reports on recent trends in the preparation of FSL batches.  The physical limits of the Parr laboratory reactor for safe  operation were also incorporated in the calculations.  The "standard" laboratory F S L batch  consisted of 360 g of IPC residue (dry basis), 700 mL of synthetic electrolyte, and 500 mL of either distilled or deionized water to simulate the steam condensate formed in the plant. The choice for synthetic electrolyte was based on expected undesirable aging phenomena with plant electrolyte, such as precipitation and degradation of organic additives. Given a density of 1230 kg/m for laboratory electrolyte containing 45 g/L copper and 3  220 g/L acid, the solids content of the "standard" slurry was 21% (wt).  Contrary to plant  Chapter 4  Experimental Work  48  practice, batch make-up was usually not modified for different IPC feed compositions in the laboratory. 4.2.3  Electrolyte preparation  Synthetic electrolyte was prepared fresh for each leaching test. It was made by dissolving the appropriate amounts of reagent-grade C u S 0 - 5 H 0 and concentrated H S 0 in distilled (Copper 4  2  2  4  Cliff) or deionized (Vancouver) water in a 1 L volumetric flask, which was diluted to volume only after cooling to room temperature. For the autoclave runs involving the use of organic additives, aqueous stock solutions of 10 g/L of both leveling agent (Crodaglu 1M44) and anti-misting agent (Dowfax 2AO) were prepared. Small quantities of both reagents were received from the C R E D plant. Crodaglu was carefully dissolved in hot water of 40-50°C, and all animal glue "solutions" were stored in a refrigerator to avoid degradation of the natural macromolecules. discarded.  After three days of storage they were  Solutions of the Dow surfactant were gently homogenized to avoid excessive  foaming. The desired quantity of organic reagents was pipetted into the electrolyte from dilute additive solutions.  Since the oxidation state +3 of arsenic is predominant at F S L conditions, reagent-grade anhydrous sodium meta-arsenite (NaAs0 ) was utilized as a source of arsenic for electrolytes 2  "spiked" with this element. It was found that As(III) can be conveniently dissolved in a neutral copper sulfate solution through the intermediate copper arsenite "Scheele's green" [59]: CuS0  + 2NaAs0  4  + xH 0  2  2  -> Cu(As0 ) 2  2  • xH 0 2  + Na S0 2  (4-1)  4  The hydrated green precipitate was obtained by mixing C u S 0 and N a A s 0 in a small volume of 4  2  water. It redissolved instantaneously upon acidification with the simultaneous release of copper and trivalent arsenic into solution: Cu(As0 ) xH 0 2  2  2  + 2H S0 2  4  -> CuS0  4  + (AsO) S0 2  4  + (2 + x)H 0 2  (4-2)  Oxidation of As(HI) to As(V) by air is negligible at room temperature because of unfavorable kinetics [47].  Let it also be noted that, contrary to oxidative sodium-iron-sulfate-containing  leaching systems, the precipitation of sodium jarosite is not of concern under the metathetic conditions prevalent in the FSL process.  Chapter 4  Experimental Work  4.2.4  Equipment  4.2.4.1  Autoclaves  49  At the C R E D laboratory a standard 2 L all-titanium Parr autoclave with Teflon liner was used. Temperature control was achieved manually both by adjusting the amount of electric current admitted to the heating mantle with a transformer, and by regulating the flow of cooling water through serpentine coils.  The presence of the poorly heat-conducting Teflon body in the  autoclave bomb greatly complicated this manual operation due to build-up of residual heat between the titanium container wall and the liner. Not surprisingly, the liner gradually deformed in the course of the preliminary autoclave experiments as a result of localized "hot spots" and thermoshock phenomena upon introduction of water to the cooling coils.  During test #13 it  became so severely dented that it obstructed the impeller. Thus, a Pyrex glass liner was used for the leaching program in Vancouver. Given the occurrence of reducing FSL conditions due to copper depletion, titanium is most certainly not the preferred autoclave construction material. Therefore, several Carpenter and Hastelloy alloys were considered for the pressure leaching program at U B C . Since these were all too expensive, stainless steel (SS) type 316 ultimately became the construction material of choice.  The 2 L autoclave in Vancouver was assembled from an assortment of custom-made and old and new Parr parts with optimum suitability for laboratory first-stage leaching in mind. Figure 4-2 shows the complete reactor set-up. The reactor was fitted with a magnetic agitator drive. The stirring shaft - the only non-SS part of the reactor internals - was made of highly resistant Hastelloy-C. It was held in place by two stirrer brackets, and two 6-bladed, axial-flow (45° pitch, down-draft) propellers were attached to it. An important safety feature of the autoclave was the SS flexible high-pressure hose attached to the head of the bomb. In case of a blow-out through failure of the rupture disc, this hose would guide the flashing slurry into the "autoclave safety vessel" filled with water.  Chapter 4  Experimental Work  Figure 4-2.  50  Autoclave set-up for the leaching program at U B C .  It was found that the fit o f Parr manufactured serpentine cooling coils into the glass liner was too tight.  Therefore, single thick-walled SS coils machined at the shop o f the Department were  utilized instead.  Despite the presence o f the Pyrex liner i n the reactor bomb, excellent  temperature control (±2°C) was achieved with the closed feed-back system comprising the Parr 4842 controller, thermocouple, electronically operated cooling water valve and coil. A i r cooling with ascending air between the container and the heating mantle proved to be unnecessary.  Thanks to thorough cleaning o f the autoclave internals after each experiment, visual inspection for excessive corrosion and wear, preventive maintenance and periodic pressure tests with water for leak detection, the SS pressure reactor has performed satisfactorily throughout the entire leaching campaign. Nevertheless, corrosion was severe, and all agitator parts have been replaced twice in the course o f the test work.  4.2.4.2 V a c u u m desiccators  A s all residues dealt with during this thesis project were prone to oxidation, reliable operation o f vacuum desiccators was essential.  The desiccators utilized in Copper C l i f f were simply  Chapter 4  51  Experimental Work  connected to the general vacuum system of the C R E D plant. In the absence of such utility in the laboratory in Vancouver, special provisions had to be made. Figure 4-3 shows the complete vacuum desiccator set-up for drying of methanol-treated residues and thawing of IPC samples at UBC.  Figure 4-3.  Vacuum desiccator set-up for the leaching program at U B C .  Two Vaseline-greased Pyrex desiccators were connected to a powerful rotary vacuum pump via a sorption system consisting of a washing bottle and two U-tubes in series.  Since all vapors  released by residues inside the desiccators passed through the pump's oil reservoir (2.5 L capacity), the drying tubes were implemented to avoid pump damage by fouling of the lubricant with condensation products from the gaseous phase. Poisoning of the oil by methanol was feared to be particularly harmful. Thus, the washing bottle was put into place to lower the concentration of methanol within the vapor phase and decrease the overall vapor pressure at the same time. The sorption system was equipped with several stopcocks and quick connectors to regulate build-up and release of vacuum as desired. Silica gel from the desiccators and U-tubes was periodically regenerated by heating and reused.  4.2.5  Autoclave leaching procedure  Since the effective volumes of the Parr autoclaves in Copper Cliff and Vancouver were practically identical, no modifications had to be made in the laboratory FSL batch make-up as described earlier.  However, minor improvements in the experimental method were brought  about throughout this project, and the autoclave leaching procedure outlined here is the finally adopted one.  Chapter 4  52  Experimental Work  The IPC residue sample was thawed overnight in its plastic bags in a vacuum desiccator to minimize atmospheric exposure. Following moisture analysis, the desired quantity of wet IPC filter cake was introduced to the glass autoclave liner and hand-stirred briefly with electrolyte and water in order to disperse the solids. Obviously, the amount of water added was corrected for the pore water volume of the cake. Using a high-impedance volt meter, the slurry oxidation-reduction potential (ORP) at ambient temperature was measured between a self-made platinum electrode and a Fisher saturated calomel electrode (SCE).  The quality of the reading of this electrode couple was  verified regularly with the aid of electrochemical set-ups of other graduate students at the laboratory. Moreover, the calomel electrode was refilled periodically with fresh saturated KC1. The liner was then cautiously lowered into the autoclave container and the bomb was assembled. The leaching experiment commenced only after heating to the desired temperature. The flow of cooling water to the agitator, temperature and pressure were repeatedly checked. At the end of the run, the electronically operated cooling water valve was opened to quench the slurry. As in the heat-up phase, the agitator was kept running during cooling. Generally, it took about 20 minutes to lower the apparent temperature of the leached pulp to «30°C. Subsequently, the autoclave container was removed from the heating mantle and immersed in a cold water bath for 10 minutes to dissipate residual heat. Following slow depressurization, the bomb was dissembled in a fume hood.  Known  quantities of a dilute sulfuric acid wash solution were utilized to recover any solids sticking to the agitator, cooling coil and bomb head in a separate beaker. During this rinsing process the solids in the liner settled, allowing measurement of the ORP of the supernatant leach solution at ambient temperature. While hand-stirring the sedimented pulp, the ORP of the leached slurry was also measured. The contents of the liner and the additional solids collected from the autoclave internals were filtered together using a large water-aspirated conical flask with Buchner funnel.  After  washing of the cake, it was repulped with wash solution, refiltered and rewashed to remove any residual entrained leach liquor. The volumes of the leach and repulp filtrates were measured. Duplicate samples of approximately 100 mL each were taken of both filtrates and submitted for assay. The mass of the "clean" wet cake was also recorded. The cake was then split into three fractions. One fraction was used for moisture analysis. It was dried in a small oven at roughly 60°C and 90% vacuum. Another fraction was repulped in 500 mL of analytical-grade methanol, filtered under vacuum and dried in a desiccator. The dry leach residue sample was distributed over two air-tight plastic containers sealed with tape. One vial was submitted for assay, while  Chapter 4  53  Experimental Work  the other one was kept as a back-up and for the S E M and X R D analyses discussed in Chapter 5. The third portion of the cake was discarded. Regarding the variety of IPC feeds utilized in the leaching program, it is emphasized here that a series of autoclave tests for the study of a particular FSL process parameter was performed with IPC residue samples of the same day, hence, of equal composition. Moreover, the effect of feed composition on extractions and leaching behavior can be compared between most series on the basis of the "standard" autoclave experiment.  Such run involved leaching of the "standard"  batch at 160°C and 820 rpm for 1 hour.  4.2.6  Releaching procedures  4.2.6.1  Releaching with sodium hydroxide  Four alkaline releaching experiments were carried out with vacuum-dried "normal" leach residue and an arsenic-enriched leach residue of the copper-depleted source, both prepared according to the autoclave leaching procedure previously described. The releaching response of the "normal" residue served as a blank. Leaching was done in a nitrogen-sparged, agitated beaker covered with a Styrofoam lid. To prevent excessive evaporation of the leach solution, the N flow was saturated with water prior to 2  introduction to the beaker by bubbling it through deionized H 0 in a gas washing bottle. The 2  sparger was a glass frit of medium porosity. Agitation was provided by an overhead stirrer rather than a stirring magnet to avoid excessive grinding of the residue particles.  A temperature of  50°C was maintained by water circulating through a thermostatic bath and the double-walled Pyrex glass beaker. The pH was kept at either 11 or 12 with the aid of a Fisher combined pH electrode and a Radiometer titrator pumping a l M N a O H solution.  Using fresh Fisher buffer solutions, the pH electrode was first calibrated at pH 7 and 10 at room temperature. The temperature dial on the titrator was used to adjust the readings to 50°C. The 300 mL starting volume of water was then heated.  Following the addition of the  desired amount of solids at 50°C, the natural pH of the slurry was recorded. The releaching experiments with "normal" residue were carried out at 20% (wt.) initial pulp density. This proved to be rather high in view of the unfavorable hydrodynamic conditions prevailing in the beaker. Subsequently, a solids concentration of 10% (wt.) was employed for the "copper-depleted" residue in order to attain better suspension of the particles.  Chapter 4  54  Experimental Work  Only after adjustment of the pH to 11 or 12, the first solution sample, marking time 0, was taken. The agitator was stopped during sample taking to allow the solids to settle, whereas the sparger was raised so as to provide a blanket of nitrogen between the atmosphere and the solution interface. A syringe equipped with a 0.22 pm disposable nylon disc filter was utilized to extract 10 mL of the alkaline supernatant solution through the sample hole in the lid. A n additional sample was taken after 30 minutes, while the 60 minutes sample was the releach filtrate. Washing of the filter cakes was conducted in a similar manner as with the autoclave tests, however, deionized water was utilized as the washing solution. After methanol treatment and drying, the whole cakes were submitted for assay.  4.2.6.2 Releaching with pregnant electrolyte Contrary to the alkaline releaching experiments performed with dry residue of the copperdepleted source, the releaching experiment with synthetic pregnant copper electrolyte involved the introduction of hot, concentrated C u S 0 - H S 0 4  depleted F S L slurry.  2  4  solution to a freshly prepared copper-  In view of the direct importance of this test to C R E D process control  (Chapter 5) it was carried out in duplicate. After quenching of the arsenic-enriched autoclave slurry to an apparent temperature of 80°C, the autoclave liner was removed from the reactor bomb and immersed in a circulating hot water bath of the same temperature.  The hot slurry was then covered with the same Styrofoam lid and  agitated and sparged as described before. The water bath was filled with small plastic balls to reduce evaporation and heat losses. Following heating of the liner contents to exactly 80°C, a solution head sample was taken according to the method outlined above. Expecting the incremental recoveries of cobalt, nickel and iron achieved during releaching to be negligible compared to the extractions of these elements attained in autoclave leaching at 160°C, it was hoped that this sample could be utilized for volume calibration. The slurry ORP was manually adjusted to and maintained at +85 m V  S C E  by pipetting  pregnant copper electrolyte into the liner. The electrolyte contained 100 g/L copper and 130 g/L acid and was kept at 65 °C in a separate water bath to prevent crystallization. The solution sample at time 0 was taken immediately after the desired ORP was reached. Additional samples were taken after 5, 15, 30, 60, 120 and 180 minutes, the latter being the filtrate sample.  After 3 hours the slurry was cooled and filtered, washed and treated with  methanol according to the usual practice.  Chapter 4 4.2.7  55  Experimental Work  Hydrogen sulfide analyses  As part of the preliminary leaching program at the CRED laboratory, it was examined if Drager tubes could be successfully used to measure the concentration of H S in the reactor freeboard 2  after a copper-depleted leach experiment. To do this, the autoclave pressure release valve was fitted with tubing to connect it to a Drager tube inserted into a hand-held pump. The measurements proved to be unpractical and unsatisfactory for three reasons. First of all, the choice for a Drager tube with a H S concentration range of either 0.5-20 or 20-200 ppm had to be 2  made prior to the actual determination. Secondly, the tubes could not be used at temperatures above 40°C.  This implied that the autoclave slurry had to be cooled down almost to ambient  temperature, causing a significant shift in the equilibrium: H S(g) 2  ^  H S(aq)  (4-3)  2  Finally, the amount of gas present in the autoclave plenum was not always sufficient for the completion of the prefixed number of pump strokes necessary for the H S analysis. This led to 2  unreliable concentration measurements on the one hand, and undesirable underpressurization of the autoclave bomb on the other hand. In view of these difficulties the Drager tube measurements were soon abandoned. Nevertheless, the presence of H S in the reactor freeboard under severe copper-depleted FSL conditions was 2  clearly demonstrated. Several other ideas for quantitative H S analysis were put forward, but 2  none of those was ever pursued. 4.3  Sample and Waste Management  Every sample taken in the course of the leaching program was registered in the laboratory sample log.  When shipped to Copper Cliff for assay, the samples were labeled and packaged according  to U N regulations. A small number of solutions and residues were analyzed at the International Plasma Laboratory (IPL) in Vancouver. All F S L solutions were collected in two 50 gallon drums.  They were intended for  processing at CRED. Unfortunately, the barrels became contaminated with other waste and will now be discarded via U B C authorities instead. Methanol waste was accumulated in large glass bottles. A distillation column will be set up to separate an impure methanol fraction from aqueous bottoms containing the majority of the  Chapter 4  Experimental Work  (base) metal values.  56  The distillate will be disposed of as simple organic waste, whereas the  bottoms will be mixed with the FSL solutions. Leach residue rejects and IPC residue leftovers were stored in a bucket. Water was added periodically to prevent drying of the solids, because they are mildly pyrophoric.  Although  methods for elimination of this solid waste have not been looked into in detail yet, it will likely be shipped to the C R E D plant.  Chapter 5  Results and  57  Discussion CHAPTER 5  RESULTS AND DISCUSSION  5.1  Introduction  After a brief discussion of the chemical composition of the IPC residue samples utilized in the experimental program, this chapter describes the effect of various physicomechanical and chemical parameters on the impurity extractions achieved and leaching phenomena observed during first-stage leaching. On the basis of the experimental results and the thermodynamics of Chapter 3 a tentative reaction model is then proposed for first-stage leaching after copper depletion. Subsequently, an entire section is devoted to the behavior of arsenic under copperdeficient FSL conditions and releaching studies of residues enriched with this element. Finally, an attempt is made to quantify the effect of air discharging of first-stage autoclaves. Scanning electron microscopy (SEM) and X-ray diffraction (XRD) work is integrated with the results of the leaching program. Although no metallurgical balances are written out here, it is pointed out that, in conformity with common metallurgical practice, impurity recoveries attained are reported both on the basis of the element's assay head grade as well as its calculated head grade. The latter is considered to be more representative of a given leaching system as it is computed using both solution and residue assays. The comparison of assay head and calculated head extractions, of course, is essential to recognize systematic errors in an experimental procedure or to identify questionable chemical analyses. From the worksheets attached in Appendix C it can be seen that, despite the substantial absolute and relative errors between the two grades for many tests, the respective nickel, cobalt and iron recoveries calculated on the basis of either grade generally do not differ more than « 2 . 5 % absolute. Such a difference is certainly acceptable in view of (1) the high concentration of these metals encountered in first-stage leaching, and (2) the limited accuracy of the Inductively Coupled Plasma (ICP) technique, the analytical method utilized with most samples. Except for the arsenic series, greater absolute errors are tolerated with arsenic considering its relatively low concentration amid the other impurities and copper. The more pronounced discrepancy in many of the iron balances can be attributed to autoclave corrosion, hydrolysis after long-term sample storage, or even systematically incorrect analyses. The relative difference between the assay head and calculated head grade of copper is usually well within 5%.  Chapter 5  Results and  58  Discussion  Without exception, the extractions tabulated or shown graphically in this chapter are based on the calculated head grade. Disregarding a few isolated cases, serious deviation of the arbitrarily set threshold error of 2.5% for cobalt, nickel and iron recoveries for a series of autoclave tests is mentioned in the discussion of the experimental results. It is also noted that the metal and acid concentrations of the leach filtrates are backcalculated values, computed from mass balances that incorporate the assays of the total leach filtrate (leach filtrate plus wash water) and the repulpfiltrate,and the moisture content of the wet leach residue. Since some oxidation of the wet FSL cakes may have taken place during repulping and refiltering, the reported residual copper concentrations of the leach filtrates can be slightly inflated.  This means that the final copper levels cannot be unambiguously related to the  measured ORP's. 5.2  Sample Composition  The minimum, maximum and average grades of the main constituents of the IPC residue samples obtained during the sampling program at INCO's C R E D plant in May, 1994, are shown in Table 5-1.  The "balance" is made up of a host of elements which were not studied in this project,  including other base metals such as lead, alkali metals such as magnesium and calcium, precious and platinum group metals, and selenium and tellurium.  Table 5-1.  Minimum, maximum and average grades of the main constituents of IPC residue samples taken from May 9 to 24, 1994.  O  balance  12.50  4.88  0.55  59.22  17.05  7.94  2.71  56.17  14.69  6.42  1.37  Co  Ni  Fe  As  Cu  min. grade, %  6.06  7.44  4.03  0.67  53.59  max. grade, %  7.61  9.99  5.39  0.79  avg. grade, %  7.13  8.79  4.70  0.73  element  S  tnt  The heterogeneous nature of the feed to the C R E D plant also becomes apparent from the histograms of Figure 5-1, which demonstrate the variability in the cobalt, nickel, iron and arsenic grades of IPC residue during the 16-day sampling period. Obviously, such daily changes in the feed composition are highly undesirable for the batch leaching process currently operated.  Chapter 5  Results and  % iron (wt.)  (C) Figure 5-1.  59  Discussion  % arsenic (wt.)  (D)  Variability in the grades of the principal impurity elements in IPC residue samples of May 9-24, 1994: (A) cobalt, (B) nickel, (C) iron and (D) arsenic.  5.3  Effect of Agitation and Pulp Density  5.3.1  Agitation series  The agitation series comprised four autoclave experiments of 1 hour at stirring rates of 600, 725, 850 and 1000 rpm, using the "standard" batch make-up with IPC residue of May 18.  A  frequency of 600 rpm was thought to be the lowest possible for effective particle suspension, whereas the maximum of 1000 rpm was dictated by the power of the agitator motor and available pulleys. The impurity extractions obtained are shown in Figure 5-2.  Chapter 5  Results and  60  Discussion  oo 100% <  0% 625 Figure 5-2.  750 875 agitation frequency, rpm  1000  Effect of agitation rate on the extraction of cobalt, nickel, iron and arsenic.  Figure 5-2 displays virtually horizontal lines representing « 9 5 % extraction each of nickel and iron and 90% cobalt, and great variation in the recovery of arsenic.  The erratic arsenic  extractions are not due to the different agitation intensities of the slurry, but rather are the result of the level of residual copper ion in solution at the end of the autoclave test. Although no reliable slurry ORP readings are available, it is evident that all tests but #20 at 875 rpm ended under mildly or strongly reducing conditions, causing partial or quantitative removal of arsenic from solution. Indeed, copper depletion frequently occurred throughout the leaching campaign.  As  emphasized in Chapter 4, however, batch make-up was not a variable in the present investigations. Figure 5-2 indicates that the FSL process is insensitive to the agitation rate as long as good particle suspension is provided. It can be speculated that the explanation for this finding lies in the skeletal morphology of the carbonylation residue: Figure 5-3 shows the typical "honeycomb" structure of IPC residue grains. Leaching of the finely dispersed impurity elements from the IPC copper sulfide matrix involves both intergrain and intragrain diffusion. Excluding any chemical effects for now, it is plausible that the movement of ions through the bulk of the solution is much faster than the diffusion of protons, cupric ion and impurity ions through the stagnant solution in the voids of the residue. In other words, more vigorous agitation of FSL slurry may increase the diffusion rate in the bulk solution ("macroscopic" diffusion), but the overall leaching rate remains limited by the movement of ions within the pores of the solid particles ("microscopic" diffusion).  Chapter 5  Figure 5-3.  Results and Discussion  61  S E M photograph showing "honeycomb" structures o f different porosity (1500x magnification).  The "topography" depicted in Figure 5-3 hampers quantitative analysis o f individual IPC or F S L residue grains by the energy-dispersive X-ray ( E D X ) technique. O n the one hand, X-rays emitted from valleys and hollows may not be able to reach the detector. O n the other hand, high-energy X-rays may penetrate a thick barrier while X-rays from lighter elements are blocked. The result in both instances is a distorted element concentration measurement. A n intermediate stirring frequency o f 820 rpm - on the Parr controller display - was chosen for all autoclave tests. The autoclave controller reading was verified with a tachometer and was less than 5% off the true rate o f about 860 rpm. From previous test work it is known that at 820 rpm all second-stage gas-liquid mass transfer limitations are overcome. Although the F S L system is not the same, it was felt that this could be relevant to the study o f the behavior o f arsenic in the presence o f H-,S. Furthermore, it was found that the autoclave runs smoothly at such agitation rate and does not vibrate or shake too much.  Chapter 5 5.3.2  62  Results and Discussion  Pulp Density series  The pulp density series consisted of five runs of 60 minutes at 5.0, 10.0, 15.0, 20.9 and 29.5% (wt.) slurry solids content. They were all carried out with IPC residue of May 22. Respecting the maximum allowable slurry volume of the Parr reactor for safe operation, the pulp density was changed by varying the amount of water added while keeping the electrolyte : residue mass ratio constant at 2.39. The impurity extractions obtained are shown in Figure 5-4.  •g 20% 03 O)  O  o/  0  I  5%  1  10%  1  15%  1  20%  1  25%  1  30%  pulp density, % solids (wt.) Figure 5-4.  Effect of pulp density on the extraction of cobalt, nickel, iron and arsenic.  Figure 5-4 shows that the FSL process is essentially unaffected by changes in the slurry solids content.  In terms of the earlier diffusion concept this indicates that bulk diffusion is not  influenced by solids concentrations of 5 to 30% (wt.). There is no obvious reason as to why the arsenic extractions peak at around 50 to 60%. On the basis of the present results, two scenarios for first-stage leaching of arsenic can be postulated: (1) the extraction of arsenic is kinetically slower than the solubilization of the other impurities, (2) the element is distributed over two or more separate minerals, at least one of which leaches favorably compared to the other(s) that is (are) refractory towards dissolution in the C u S 0 4  H S 0 medium. 2  4  The second scenario seems more likely. For instance, one could speculate that arsenic exists both in solid solution in the copper sulfide matrix as well as in a discrete nanocrystalline phase such as an arsenide. However, using S E M or X R D it is already a near impossible task to identify a single arsenic compound in IPC or FSL residue, let alone distinguish between two or more of them.  Chapter 5  Results and  63  Discussion  In test #27 at 30% solids - an impossible pulp density in present plant practice - an anomalously high arsenic extraction was achieved: 69.6%. The dried residue utilized in that particular test must have been very oxidized considering the low copper and high acid consumptions of 16.3 and 299.4 kg/t IPC residue, respectively. The effect of grinding or attritioning action on the impurity extractions at the higher solids concentrations is hard to discern, however, particle reduction by such mechanisms is unwanted with regard to the filtration characteristics of FSL slurry.  5.4  Effect of Temperature and Residence Time  5.4.1  Kinetic studies on IPC residue with a high Cu : S  tot  ratio  In older INCO correspondence, second-stage leaching difficulties were sometimes attributed to the action of the mineral cuprite, the presence of which in IPC residue with a high Cu : S  tot  mass  ratio can be explained as follows. With insufficient sulfur available to accommodate all of the copper as sulfides in the IPC reactor feed, the remaining metallic (nickel) phase becomes richer in copper. Therefore, after selective removal of the nickel by carbonylation, the IPC residue contains more "exposed" copper metallics. These fine copper metal particles are very amenable to oxidation to cuprite during wet grinding, transport and storage prior to first-stage leaching. The FSL process supposedly left C u 0 intact. It was postulated that the oxide locally 2  rendered the surface of the chalcocite to be leached more alkaline, which led to the formation of an adhering coating of BCS that impeded further oxidation of the sulfides. As outlined in Chapter 1, it is now known that high-(Cu : S) feeds to the SSL process can result in elevated concentrations of BCS, increased slurry viscosities and slower gas-liquid mass transfer in the reactor. With both the obsolete and the recent theory in mind, a comprehensive leaching study of high-(Cu : S ) IPC residue was undertaken. In addition, the mineralogy of various FSL residues tot  produced in interruptive and extended autoclave tests was investigated by X R D . Nineteen individual autoclave runs at 80, 120, 140 and 160°C were performed with IPC feed material from May 11-14 using the "standard" batch make-up. The Cu : S  t t  ratio for each  of these residues was around 4.3, while their overall chemical composition was very similar. Grinding data from the CCNR indicated that the grind size of IPC residue from May 11,12 and 14 was « 7 0 % passing 325 mesh, whereas residue of May 13 was coarser at 86% passing the same screen size. CRED  In view of the buffer capacity of the IPC thickener and surge tanks at the  facility, however, these data are unlikely to represent the true fineness of the  Chapter 5  64  Results and Discussion  carbonylation residue samples of May 11-14. No screening analyses were performed on these and any other IPC residues, since grind size is not a process parameter that can be easily controlled or modified at the leaching plant. The extraction curves are shown in Figure 5-5. In the interpretation of this figure it should be realized that the errors in the nickel, cobalt and iron balances for the kinetic experiments at 80 and 120°C are greater than for any other test.  After re-analysis of all filtrate samples, the  consistency of the discrepancy between assay head and calculated head extractions seems to point to incorrect concentrations of the three metals in either the IPC or the leach residues. Since assays for the head samples submitted during the sampling campaign in Copper Cliff have proven very reliable, the FSL residue analyses are believed to be inaccurate. However, due to time and budget constraints this has not been verified. No other possible sources of error have been identified. The experimental procedure was not changed. Since no anomalous increase in the chromium content of the F S L filtrates was observed, excessive corrosion of the SS autoclave internals is also ruled out; one would rather expect an inhibiting action by copper as a result of the higher concentrations of this metal in FSL slurries at lower leaching temperatures.  Analyses of the methanol filtrates showed that the  coloring of those solutions was due to negligible concentrations of impurity elements that could not possibly have been responsible for the great biases in the metallurgical balances. In Figure 5-5 only cobalt and iron follow the expected temperature-residence time trend. Long retention times at a temperature of 160°C without replenishment of the copper level in solution are detrimental to the arsenic recovery, and, interestingly, also appear to adversely affect the nickel extraction. Reducing conditions at the end of 4-hour leaching test #28 were indicated by the 35 psi (H S) pressure increase recorded by the gauge. During the same experiment a bronze2  colored precipitate was deposited on the stirring shaft, which unfortunately could not be scraped off for S E M and X R D analysis. Figure 5-5 implies that in the plant large quantities of cobalt, nickel, iron and arsenic are solubilized from high-(Cu : S ) IPC residue during first-stage batch make-up and preheating to tot  80°C and batch heating to 160°C. In fact, the vast majority of "leachable" arsenic values has already been extracted once the current FSL operating temperature is reached. The iron and nickel recoveries achieved after heating-up to 160°C are as high as 78.3 and 76.2%, respectively, clearly suggesting the presence of these metals in oxide phases. The dissolution of cobalt is more dependent on metathesis reactions, which are known to require autoclave rather than atmospheric leaching conditions to take place at industrially acceptable rates.  Chapter 5  Results and  Discussion  65  (C) Figure 5-5.  (D)  Fraction of metal extracted from IPC residues with an average Cu : S  tot  mass  ratio of 4.3, as a function of leaching time: (A) cobalt, (B) nickel, (C) iron and (D) arsenic. Note the different extraction scale in the arsenic graph.  In early INCO studies [60] it was established that the overall first-stage leaching rate is first order with respect to the residual concentration c of impurities in the leach residue:  .L±  at  =K (c-c„) T  (5-1)  Chapter 5  Results and  In equation (5-1)  66  Discussion  represents the "unleachable" impurity content of IPC residue, and K  T  is the  apparent overall rate constant. Bearing in mind the changes in the composition of IPC residue over the twenty years the C R E D plant has been in operation, the extraction curves shown in Figure 5-5 do not follow equation (5-1) at all. One could argue that the experimental method employed in this leaching campaign has failed to provide useful data for the derivation of a kinetic model describing first-stage leaching of IPC residue with a high Cu : S  tot  mass ratio within the temperature range investigated. Indeed,  in the present study the most important part of the kinetic information is lost by the time the first data are generated; the slopes of the cobalt, nickel and iron extraction curves merely reveal that the leaching rate is largely independent of temperature after heating-up to the desired leaching temperature. Moreover, the similar slopes make the calculation of apparent activation energies from Arrhenius plots impossible. One could also challenge the usefulness of the concept of the overall rate constant with respect to first-stage leaching.  The FSL process is a highly complex system that comprises  simultaneous leaching of various impurity metals from a multi-phase material according to at least three different mechanisms: (1) oxide dissolution, (2) metathesis, and (3) cementation. The system is further complicated by electrochemical interactions such as galvanic conversion. Therefore, even if apparent rate constants and activation energies could have been computed, it would have been extremely hard to assign any specific physical meaning to them. Although it is recognized at this point that leaching experiments with pure minerals would certainly give a better insight into the kinetic contributions of individual phases, it must be emphasized that IPC residue is not a naturally occurring material and has a unique composition and morphology. X R D studies on IPC and leach residues of the high-(Cu : S ) kinetic series were performed at tot  Geological Sciences at U B C using a Siemens D-5000 diffractometer with Diffrac/AT software. Prior to taking a spatula tip of residue sample, the sample vial was thoroughly shaken. The small quantity of residue was then wet-ground in methanol in an agate mortar, quickly dried under a hot lamp, and transferred onto a glass slide using wax paper. After cutting of the sample with a razor blade, several drops of methanol were added to allow the solids to be smeared out evenly. Once the methanol had evaporated again the slide was inserted into the sample holder of the diffractometer. IPC residue samples of May 12 and 13 were first scanned from 20 = 5 to 75° in steps of 0.02° with a counting time of 1 second. On the basis of those patterns it was decided to reduce the 29 scale to 28-58° as all major peaks are located within that range, while increasing the  Chapter 5  Results and Discussion  67  counting time to 2.5 seconds to improve the peak : noise ratio.  Following calculation and  subtraction of the background noise, all patterns were smoothed. The X R D patterns for the leaching tests at 160°C with IPC residue of May 12 are shown in Figure 5-6.  30  32  Figure 5-6.  Stacked X R D patterns showing mineralogical changes in the leach residue over time at 160°C.  Key: ch = chalcocite, cp = cuprite, cu = copper,  dj = djurleite, pi = pentlandite and tn = tenorite.  As anticipated, elemental copper is one of the major phases present in the IPC and leach residues of Figure 5-6. The shoulder in the main copper peak (29 = 43.3°; d = 2.09 A) as well as other smaller peaks are evidence for the presence of troilite (FeS). The high intensity of the principal copper peak may be due in part to the presence of minor amounts of iron-rich trevorite, (Fe,Ni)Fe 0 , or, more likely, bunsenite (NiO), to which the mysterious peak at 20 = 37.2° 2  4  (d = 2.42 A) could possibly also be ascribed.  This peak might also belong to another, still  unidentified (synthetic) mineral phase. Chalcocite (Cu, S ) is the main copper sulfide mineral in the IPC feed, but djurleite 96  (Cu,  93  S ) is formed during leaching.  Tenorite (CuO) and cuprite (Cu 0) are the chief oxide 2  phases, yet X R D is inconclusive as to whether the latter was really formed prior to lixiviation in the C u S 0 - H S 0 medium or just through handling of the leach residues. 4  2  4  Chapter 5  Results and  68  Discussion  From the extraction curves in Figure 5-5 it can be deduced that the removal of arsenic from solution after several hours of leaching at 160°C was initiated by a persistant demand for an oxidant for the leaching of a residual cobalt- and iron-bearing phase. As can be seen from Figure 5-6, there is indeed less cobaltiferous pentlandite ((Co,Ni,Fe) S ) present in the leach residue 9  after 4 hours than after 1 hour of leaching.  8  Unfortunately, no arsenic minerals could be  identified. The mineralogical composition of the other high-(Cu : S ) IPC feed materials was almost tot  identical, while the inversely proportional relationship between impurity extraction and peak intensity was evident. 5.4.2  Kinetic studies on IPC residue with a low C u : S  t o t  ratio  For comparison to the results reported in section 5.4.1, kinetic experiments at 160°C on IPC residue with a Cu : S  tot  ratio of only 3.2 were also performed. To avoid the highly reducing  conditions encountered in the course of test #28, 800 mL of electrolyte instead of 700 mL was utilized in the batch make-up for these series. Four-hour autoclave run #49 had to be stopped prematurely because of agitation problems. Although the sulfide minerals in IPC residue of May 19 matched those present in the high(Cu : S ) feeds, the oxide mineralization consisted entirely of spinel-type phase(s) and tot  bunsenite; no evidence for the presence of either cuprite or tenorite was found in the diffractogram.  Surprisingly, elemental copper was also a component of the low-(Cu : S ) tot  residue, although the intensity of its peaks was somewhat lower.  The extraction curves obtained are shown in Figure 5-7. The low-(Cu : S ) material proved to tot  be highly reactive, and the increased concentration of cupric ion in the slurry was not quite enough to eliminate copper depletion. The maximum arsenic extraction of 59.5% was attained only after 30 minutes of leaching. Again the nickel and iron recoveries were near 80% at time zero, whereas the metathesis reactions for the solubilization of cobalt took longer to reach such a level.  Chapter 5  Results and  Figure 5-7.  69  Discussion  Fraction of metal extracted from IPC residue with a Cu : S  tot  mass ratio of  3.2, as a function of leaching time. 5.5  Effect of Electrolyte Composition  5.5.1  C o p p e r series  The copper series comprised five 1-hour leaching tests with electrolyte copper concentrations of 60, 45, 30, 15 and 0 g/L, together with 220 g/L H S 0 . Unlike all other experiments discussed in 2  4  this chapter they were carried out at the C R E D laboratory. Each batch was composed of 360 g IPC residue of May 20, 700 mL of electrolyte and 500 mL deionized water.  From the pulp  density series it follows that the changes in the slurry solids concentration caused by the different electrolyte copper levels have had negligible impact on the leaching results of the copper series.  In the course of the preheating phase of test #13 at 0 g/L copper, the Teflon liner employed in Copper Cliff became so severely deformed that it blocked the agitator. The autoclave run was discontinued and never repeated since its academic value was found not to weigh up to the risk of equipment damage due to the large quantity of H S (and possibly H ) formed under zero copper 2  conditions.  2  Indeed, at initial electrolyte copper contents of 30 and 15 g/L the reactor gauge  displayed pressure increases of 25 and 160 psi, respectively, from the normal 80 psi.  The  presence of hydrogen sulfide gas in the freeboard for those tests was demonstrated with the aid of Drager tubes. The formation of H S is also reflected by the steady increase in acid consumption 2  with decreasing copper concentration.  Chapter 5  Results and  70  Discussion  The copper series leaching results are plotted in Figure 5-8. en  100% r  copper concentration in electrolyte, g/L Figure 5-8.  Effect of electrolyte copper concentration on the extraction of cobalt, nickel, iron and arsenic.  Figure 5-8 shows that the iron extraction is barely influenced by the level of cupric ion in the electrolyte, which supports previous comments on the presence of this metal in IPC residue oxide minerals. The consistently high cobalt recoveries are due to metathesis of this metal with arsenic or acid once copper has been depleted from solution. At the low-copper side it is not clear if arsenic is leached in the first place or if it is "reprecipitated" at a later stage. In the case of severely reducing conditions, dissolved nickel is also removed from solution. The decrease in the arsenic and nickel recovery is 52.4 and 12.1%, respectively, when the copper content of the electrolyte is lowered from 60 to 15 g/L. In contrast, a large excess of copper ions seems to have a beneficial effect on the arsenic extraction, yet no duplicate test was carried out to assess the reproducibility of the results of test #9 at 60 g/L. Strikingly, the leach residues obtained with low-copper electrolyte are lighter than the ones produced when electrolyte with an elevated copper level is utilized.  This finding can be  attributed to the lower grade of "heavy" copper of the former.  5.5.2  A c i d series  The acid series consisted of five leaching tests with electrolyte acid concentrations of 0, 50, 100, 150 and 220 g/L, while the copper level was maintained at 45 g/L. IPC residue of May 9 was chosen for these series because of its high oxygen content of about 8%.  As with the copper  Chapter 5  Results and  71  Discussion  series, the variation in pulp density arising from the changing acid content of the electrolyte in each experiment can be ignored. The impurity extractions obtained are shown in Figure 5-9. (o 100%  <  50  100  150  200  250  acid concentration in electrolyte, g/L Figure 5-9.  Effect of electrolyte sulfuric acid concentration on the extraction of cobalt, nickel, iron and arsenic.  The dependency of iron extractions on the electrolyte acid concentration is further substantiated in Figure 5-9. Since the cobalt extractions achieved do not differ more than approximately 20% over the entire range of acid levels investigated, it can be concluded that this metal is solubilized from May 9 residue predominantly by metathesis reactions. This was really already evident from the color of the leach filtrates, which changed from deep red to the usual dark brown with increasing H S 0 levels in the electrolyte. 2  4  Unfortunately, the acid series results for arsenic and nickel are masked by copper depletion. Arsenic is likely "reprecipitated" quantitatively and, although a link between the nickel recovery and the availability of acid is established, the extractions of this metal are biased by partial reprecipitation. At the regular electrolyte acid concentration of 220 g/L its recovery is some 5% lower than the extractions of iron and cobalt.  Nevertheless, Figure 5-9 demonstrates that the sensitivity of the impurities to the amount of acid in solution decreases in the order Fe > Ni > Co, which is believed to be true for first-stage leaching in general.  For all three metals the extraction is almost linear with the acid  concentration until it levels off at 100 g/L.  Considering the spinel-type structure of most  impurity oxides in IPC residue, in a broader sense this behavior is analogous to the leaching of franklinite (ZnFe 0 ) in the zinc industry. Although leaching of this zinc spinel is practiced at 2  4  Chapter 5  Results and  72  Discussion  atmospheric conditions in a Z n S 0 - H S 0 4  2  4  lixiviant, its break-up is also strongly affected by  acidity. Test #14 at 0 g/L acid produced the sole leach residue in the entire leaching program heavier than 360 g, the mass of IPC residue used in most batches.  Strangely, X R D studies showed no  evidence for the presence of copper hydrolysis products or basic nickel or cobalt sulfates in the dry cake. This means that its large mass is due to a combination of incomplete dissolution of oxide minerals, as indicated by its high oxygen content and low Cu : O mass ratio, and the replacement of lighter base metals (cobalt) by heavier copper through metathesis reactions. It is also worth noting that in the same experiment a minor quantity of sulfuric acid was produced. Although high impurity extractions can be obtained at 150 g/L, the acid surplus is certainly necessary to prevent massive hydrolysis of dissolved metals during cooling in general and filtration in particular. 5.5.3  Additives series  Originally, two autoclave experiments with IPC residue of May 10 were planned for the additives series: a blank run without organic reagents, and a "worst case" run with 5 ppm Crodaglu 1M44 and 100 ppm Dowfax 2A0 in the electrolyte. Utilizing the usual 700 mL of electrolyte, the leach conditions became so reducing at the end of both tests that it was decided to repeat them with 800 mL instead.  When the slurry ORP's measured upon completion of the additional  experiments were still negative, the additives series was ceased since higher priority was given to other test work.  The leaching results, which are not less informative, are summarized in  Table 5-2.  As anticipated, the presence of additives in the electrolyte has no impact on first-stage leaching, even at the uncommonly high levels used in tests #48 and #55.  Crodaglu degrades rapidly at  temperatures higher than 60-70°C, so that glue essentially decomposes during preheating of the batch. Although Dowfax has been attributed good thermal stability and resistance to deactivation by other ions even in concentrated electrolytes [61], the combination of the high leaching temperature and the harsh FSL solution is likely detrimental to Dowfax's di-phenyl oxide disulfonate structure. Moreover, the surfactant acts at the solution/gas interface rather than at the solution/solids interface.  It is noted, however, that minor quantities of elemental sulfur-like  flakes were observed floating on the supernatant solution during filtration, yet assays for S° in the additives residues were zero.  Chapter 5  Results and  Table 5-2.  73  Discussion  Effect of organic electrowinning additives on the extraction of cobalt, nickel, iron and arsenic.  electrolyte volume  700 mL  800 mL  5.6  impurity extractions  impurity extractions  no additives added  additives added  93.6% Co  93.8% Co  87.2% N i  87.9% N i  0.0% As  0.0% As  96.2% Fe  95.8% Fe  93.4% Co  93.5% Co  92.2% Ni  94.5% N i  0.0% As  0.0% As  95.4% Fe  95.8% Fe  Reaction Model for First-Stage Leaching under Reducing Conditions  Interestingly, the results of the additives series suggest that 95% nickel extractions can still be obtained at a slurry potential as low as -263 n i V  S C E  at room temperature, when all copper and  arsenic have already been depleted. In addition to the results of the agitation series, this is strong evidence for a sequence of "precipitation" reactions rather than "co-precipitation" of arsenic and nickel after copper depletion.  On the basis of the experimental results so far, it is postulated that dissolved arsenic takes over the role of copper in the metathesis reactions once the copper level in the FSL solution has been reduced to zero. Since the slurry ORP is proportional to the logarithm of the activity of dissolved species, a major shift in the slurry potential is observed whenever the last traces of a certain metal are removed from solution and another redox system becomes the controlling one.  With the  successive depletion of copper and arsenic, the FSL moves to progressively lower potentials.  From the E - p H diagrams in Chapter 3 it follows that the aforementioned slurry ORP of h  -263 n i V test #55.  S C E  is well within the stability region of H S at the residual filtrate acidity of 95.2 g/L of 2  It is believed that the removal of arsenic from solution is dominated by exchange  reactions with unleached sulfides until the point of hydrogen sulfide formation is reached and the element is readily precipitated.  Chapter 5  Results and  74  Discussion  After arsenic, nickel also precipitates according to: Ni  2+  + HS 2  -> NiS +  (5-2)  2H  +  but this reaction is both thermodynamically and kinetically much less favorable.  Using  thermodynamic data from Appendix A , an equilibrium constant of only 82.4 at 25°C can be calculated for reaction (5-2). The observed build-up of H S in the autoclave freeboard and the 2  decreased filterability of reducing F S L slurries could both be indicative of the fact that the precipitation of millerite is homogeneous. There is one more scenario regarding the behavior of arsenic and nickel that must be considered. Under the right conditions, two co-existing sulfide minerals will react to form a mixed sulfide phase with a more negative Gibbs free energy (greater stability) than obtained by the sum of the Gibbs values of the individual minerals (after E. Peters). Thermodynamically speaking, the conditions for the formation of gersdorffite or wolfachite (both NiAsS) are quite favorable.  First of all, when Figures 3-3 and 3-5 are  superimposed a significant overlap of the arsenic and nickel sulfide predominance areas becomes apparent.  Secondly, a strong reductant R (metallic copper) is present in the F S L system,  allowing reactions of the type: 2 NiS + As S 2  + 6R+6H ^>2 +  3  NiAsS + 6 R  +  + 3 HS 2  (5-3)  From a kinetic point of view, it remains to be determined at what rate reaction (5-3) would proceed at the FSL temperature of 160°C. 5.7  Behavior of Arsenic  5.7.1  A r s e n i c series  The aim of the arsenic series was to gain better insight into the relationship between end-of-leach slurry potentials and reprecipitation of arsenic. Furthermore, it was hoped that the series would produce FSL residues more suitable for SEM and X R D work.  The arsenic series comprised six autoclave runs of 0, 10, 20, 30, 60 and 240 minutes duration. They were carried out with IPC residue of May 24, and an electrolyte containing 35 g/L copper, 220 g/L sulfuric acid and a 10 g/L arsenic "spike". The choice for the May 24 feed was solely  Chapter 5  75  Results and Discussion  based on the availability of sample, whereas the copper level of 35 g/L was selected to achieve acceptable extractions of nickel, cobalt and iron (copper series) while deliberately inducing arsenic precipitation. During the preparation of electrolyte with 25 g/L each of copper and arsenic for additional test #66, the solubility product of either sodium arsenite or arsenic(IU) oxide was exceeded. This leaching experiment was therefore canceled.  50%  0  60  120  leaching time, min  180  240  Figure 5-10. Extraction of cobalt, nickel and iron achieved in the arsenic series, as a function of leaching time.  60  120  leaching time, min  180  Figure 5-11. Arsenic precipitation and end-of-leach slurry potentials as a function of leaching time.  Chapter 5  Results and  76  Discussion  The results of the arsenic series are shown in Figure 5-10 and Figure 5-11.  Looking at  Figure 5-10, two features are evident immediately: (1) reprecipitation of nickel, and (2) a steady increase in the recovery of cobalt and iron even when copper and arsenic are completely and nickel is partially removed from solution. The reason for the substantial rise in the extraction of cobalt (as in Figure 5-5) and iron is due to exchange reactions between unleached sulfides and acid. The behavior of nickel is in agreement with the reaction model described in the previous section. From Figures 5-10 and 5-11 it follows that after 30 minutes of leaching 94% nickel has been extracted from the IPC residue, while more than 90% of the arsenic has precipitated. Thus, arsenic precipitation is almost completed at -75 m V  S C E  at 25°C, whereas - in accordance with  Figure 3-5 - nickel removal from solution is initiated between -75 and -305 mV. The  existence of a plateau in Figure 5-11 at roughly -40 mV is unlikely.  The  discontinuity in the precipitation curve is presumably due to the small difference in the autoclave leaching times for which the data were collected. In the arsenic series, the distinction between precipitation and reprecipitation of arsenic is unnecessary. This is shown by both the invariably positive percentage of arsenic removed from solution, as well as by the value of the weight ratio As in IPC residue : As in leach residue, which never becomes significantly smaller than 1. Hence, either no arsenic at all is extracted from the feed or any dissolved arsenic from the IPC residue has already reprecipitated after heating of the batch to 160°C.  The fact that the mass of leach residue first declines with time and later on  increases again is indicative of the competition between extraction and precipitation. Several loose powder mounts and polished sections of arsenic series residues were examined by S E M with E D X . Despite the relatively high arsenic content («3%) of some of those residues, the S E M work can be considered unsuccessful. No rims of arsenic sulfides were found; instead, the element seems to be distributed quite evenly throughout the sample. This is believed to be due to the extremely high specific surface area of the "honeycombs", which offer an almost infinite number of sites for precipitation. Since lead peaks overlap part of the arsenic spectrum in E D X , mapping for arsenic was rendered virtually impossible by the presence of lead in the samples. The chemical composition and morphology of IPC and FSL residue varies significantly from grain to grain.  Dense grains are usually much smaller than particles with the porous  skeletal structure. The latter cannot be analyzed reliably by the E D X technique in view of the earlier explained topographic effects.  Chapter 5  Results and  77  Discussion  Using the most promising residue (produced in the arsenic series) in X R D studies, slow scanning around arsenic sulfide peaks did not provide the information hoped for.  No decisive phase  identification was achieved with residue #61 containing 3.2% arsenic and 2.2% nickel.  5.7.2  Releaching with sodium hydroxide  Leaching experiments and S E M and X R D work together have failed to provide physical evidence that explicitly points to the precipitation of arsenic sulfide compounds under copper-depleted FSL conditions. The only true experimental indication of the formation of such precipitates is the presence of H S in the autoclave. Thus, in the absence of direct proof for the occurrence of 2  A s S , the goal of the alkaline releaching experiments was to gather additional indirect evidence x  y  for the existence of arsenic sulfide precipitates. Time constraints have limited the number of reextraction tests carried out in a NaOH medium to the four described in this section.  In cyanidation circuits all over the world, arsenic sulfide minerals are known to be troublesome gold ore components.  Indeed, orpiment and realgar are capable of fouling caustic solutions  through the formation of arsenites and thioarsenites: As S  + 6 NaOH  -» Na As0  3  + Na AsS  3  + 3 H0  (5-4)  2 AsS + 6 NaOH  Na As0  3  + Na AsS  2  + 3 H0  (5-5)  2  3  3  3  3  3  2  2  As part of the present investigations into the behavior of arsenic, redissolution of the element from an arsenic-enriched dry FSL residue was tested at pH 11 and 12.  For comparison, the  amenability of a "normal" residue to leaching in caustic soda solutions of the same alkalinity was also tested.  The arsenic re-extractions achieved in the 1-hour experiments are shown in  Table 5-3.  Table 5-3.  Effect of residue type and alkalinity on the re-extraction of arsenic.  residue type  pH  arsenic re-extraction  low-arsenic  11  0.0%  0.53% As  12  1.9%  high-arsenic  11  11.0%  2.88% As  12  16.0%  Chapter 5  Results and Discussion  78  Contrary to the excellent balances of the arsenic series, poor arsenic accountability was attained in the alkaline releaching experiments. This is due to the low concentrations of the metalloid in the NaOH solutions. The redissolution results are disappointing and essentially inconclusive. It is believed that the low arsenic recoveries reported in Table 5-3 are caused by unfavorable experimental conditions, such as a low temperature, a short leaching time and insufficient alkalinity,' rather than by the presence of non-sulfidic arsenic compounds in the residue. This view is supported by the fact that no other elements solubilized in the high-pH solutions. The unsatisfactory arsenic extractions could equally be attributed to the conversion of one arsenic sulfide compound into another with the release of sulfide ion instead of arsenic. However, no sulfide assays are available to confirm such a theory.  Finally, the different  reactivity of wet and dry FSL residue might have played a role. 5.7.3  Releaching with pregnant electrolyte  The purpose of this important releaching experiment was to investigate the possibility of arsenic redissolution from copper-depleted FSL slurry through the introduction of pregnant electrolyte to the first-stage autoclave pressure let-down tank. Since a line for Durco Filtrate to the first-stage Product Holding Tank (PHT) could easily be put in place at the C R E D plant, this would be a convenient way to "correct" reducing batches prior to filtration and alleviate the arsenic related problems during second-stage leaching.  In the laboratory simulation of this electrolyte addition option, the mass balances over the autoclave are corrupted by unavoidable losses of leach residue to the autoclave internals. Therefore, the releaching results in Appendix C are reported as the relative distribution of arsenic over the liquid and solid phases as a function of time. It is noted that any arsenic removed from the liner through sample taking has been accounted for.  The duplicate releaching experiment with hot C u S 0 - H S 0 solution was carried out at a slurry 4  ORP of 85 m V  S C E  2  4  at 80°C. This redox potential was suggested by C R E D staff and is based on a  study at the JRGRL, which recommended such a value for optimum («50%) arsenic extraction in the FSL process. Unfortunately, the arsenic re-extractions of 1.4% and 6.2% attained after 3 hours clearly indicate that arsenic redissolution at 85 mV is unsuccessful.  In fact, this had already become  evident in the course of the test since negligible changes in the slurry ORP were observed once  Chapter 5  Results and  79  Discussion  the desired redox value was reached. Thus, similar to the concentration of other impurities in solution, the elevated copper levels of about 17.5 g/L and 10.5 g/L, respectively, remained basically constant. It is believed that the conditions at which the re-extraction experiments with synthetic Durco Filtrate were done are both thermodynamically and kinetically unfavorable.  From the  potential-pH diagrams of Figure 3-3 and 3-4 it can be deduced that the potential of 85 mV at high acidity is thermodynamically  well within the stability region of arsenic sulfide. Furthermore,  metathesis reactions generally require autoclave conditions to take place at industrially acceptable rates. Hence, the temperature of 80°C, or, the assumed temperature in the first-stage flash tank, is kinetically  disadvantageous.  Two process control options for resolubilization of arsenic remain to be examined, namely (1) releaching with pregnant electrolyte at other potentials and temperatures, and (2) intermittent electrolyte injection during first-stage leaching. With regard to the first option it is advocated the experimental method be improved by connecting the platinum electrode in the releach vessel with an external calomel reference electrode using a salt bridge. The additional potential drop so introduced is negligible, while the life  of the  reference  electrode  is  greatly prolonged as  disproportionation of the calomel is eliminated.  thermal degradation through  Moreover, the risk of unreliable potential  measurements caused by desaturation of the KC1 solution is much smaller. The second option calls for changes in batch make-up, and the installation of ORP probes and a high-pressure electrolyte injection system. Before the implementation of these changes, the incremental arsenic recovery achieved by the introduction of copper electrolyte needs to be investigated.  Obviously, intermittent electrolyte injection would be an important first step  towards a continuous leaching process. In the meantime, the installation of an ORP probe between the first-stage PHT and Filter Feed Tank seems to be the best option. On-line potential measurements would be a suitable way to identify copper-depleted batches and adapt the make-up of following batches.  5.8  Air Discharge Series  In recent years, staff at the C R E D plant have considered the use of compressed air instead of "push steam" to discharge the first-stage autoclaves. Ball park estimates for compressor capacity and cost for air discharge were made back in 1992, together with a proposal for the mode of operation [62].  Chapter 5  Results and  80  Discussion  The important disadvantage of the current practice is overheating of the leach solution, and hence, the acid-resistant lining, to temperatures as high as 175°C by "push steam" entering the FSL reactor. This makes the bricks and mortar prone to cracking and spalling upon introduction of a new batch of only 80-90°C almost directly after discharge of the hot slurry. It is believed that the use of air would circumvent this heat management problem and ultimately result in longer lining life. As mentioned in Chapter 4, the projected series of laboratory pressure leaching experiments followed by air discharge of the autoclave were dropped. The basis for this decision is the outcome of the simplified calculations attached in Appendix D. These show that air discharge of a typical FSL batch would cause an unwanted rise in the copper level in solution of at least 0.35 g/L, and a multiple thereof should the discharge pressure indeed be maintained at 115 psig. The exothermic heat effect associated with air discharge is hard to compute; it would likely be small and pose no threat to the brick lining.  In view of significant redissolution of  cupric ion, however, setting up a detailed energy balance over a FSL autoclave would be a futile exercise. Thus, it can be concluded that air discharge of the FSL reactors would result in both considerable copper losses to the tailings and a higher copper content of the mixed carbonate product. Moreover, even if the the increase in the copper concentration was acceptable, the use of air would be unsafe considering the possible presence of H S or H in the freeboard after a copper2  2  depleted leach. For the same reason, controlled air injection into the first-stage reactor has been ruled out as a method to dissolve reprecipitated arsenic. Finally, the current practice of mixing of the off gases from the F S L and SSL could be dangerous in case of severely reducing first-stage conditions caused by process upsets.  Chapter 6  Conclusions and  Recommendations  81  CHAPTER 6  CONCLUSIONS AND RECOMMENDATIONS  1)  The impurity extractions achieved in the FSL process are independent of the agitation rate and pulp density as long as residue particles are freely suspended. Although more intense stirring of FSL slurry may enhance the diffusion rate of ions in the bulk of the solution, the overall leaching rate remains limited by the movement of ions participating in the leaching reactions inside the skeletal residue structure.  2)  The sensitivity of the impurities to the amount of acid in solution decreases in the order Fe > Ni > Co, indicating that a significant portion of the iron and, to a lesser extent, nickel are present in oxide phases in the IPC feeds, whereas the dissolution of cobalt is largely due to metathesis reactions. As anticipated, the efficiency of the FSL process is not affected by the presence of organic electrowinning additives such as Crodaglu 1M44 and Dowfax 2A0 in the slurry, since these decompose at the high temperature and acidity prevailing in the first-stage reactor.  3)  Regardless of the copper level in solution at the end of the first-stage leach, the arsenic recovery generally peaks at about 50 to 60%.  On the basis of the present results, two  scenarios for first-stage leaching of the metalloid can be postulated: (1) the extraction of arsenic is kinetically slower than the solubilization of the other impurities, (2) the element is distributed over two or more separate minerals, at least one of which leaches favorably compared to the other(s) that is (are) refractory towards dissolution in the C u S 0 - H S 0 4  2  4  medium. As both S E M and X R D have not proven useful for the identification of arsenic phases in this project, a series of extended leaching experiments with intermittent electrolyte addition should be carried out to investigate both scenarios in more detail. Simultaneously, the test results would provide insight into the leaching response of the first-stage process in a continuous operation.  4)  Kinetic studies have not produced useful data for the derivation of a kinetic model describing overall first-stage leaching. However, they have shown that the leaching of IPC residue does not or no longer satisfy the first order relationship proposed by INCO in the early 1970's.  Moreover, only cobalt and iron follow the expected temperature-  retention time trend; long residence times  at a temperature of  160°C  without  Chapter 6  Conclusions and  82  Recommendations  replenishment of the copper level in solution are not only detrimental to the arsenic recovery, but, interestingly, also adversely affect the nickel extraction. On the basis of thermodynamics and the behavior of arsenic and nickel observed in the experimental work, a rudimentary model has been proposed to describe the presumably sequential processes taking place after copper depletion.  The model suggests that  dissolved arsenic takes over the role of copper in the metathetic exchange reactions with unleached sulfides once this metal has been quantitatively consumed.  As the potential  becomes progressively lower, H S becomes stable and precipitates any residual arsenic in 2  solution. In both ways arsenic reports to the leach residue as an arsenic sulfide. After arsenic, nickel is also precipitated, but the formation of millerite (NiS) is thermodynamically much less favorable as can be deduced from the small equilibrium constant at 25°C for the reaction between nickelous ion and dissolved hydrogen sulfide gas.  Furthermore, the observed accumulation of H S in the autoclave plenum and the 2  inferior filtration characteristics of severely reduced F S L slurries could point to kinetically slow homogeneous precipitation. Unfortunately, neither direct nor indirect proof of the presence of arsenic sulfides in FSL residues of the copper-depleted type has been produced. Using S E M , mapping of arsenic in residues enriched with this element was made impossible by the presence of lead. The concentration of arsenic phases is generally too low for identification by X R D . Finally, the results of releaching experiments aimed at the selective re-extraction of arsenic from leach residues in NaOH solutions are essentially inconclusive. 5)  Arsenic redissolution from a copper-depleted FSL batch cannot be accomplished by the introduction of pregnant electrolyte (Durco Filtrate) to the first-stage autoclave pressure let-down vessel under the conditions of the laboratory simulation. Two options for the re-extraction of arsenic remain to be studied, namely (1) releaching with pregnant electrolyte at higher slurry potentials than 85 m V  S C E  and  temperatures than 80°C, although the latter likely already approximates the temperature in the flash tank, and (2) intermittent electrolyte injection during first-stage leaching. In the meantime, the installation of an ORP probe between the first-stage PHT and Filter Feed Tank seems to be the best option.  On-line potential measurements would be a  suitable way to identify copper-depleted batches and adapt the make-up of following batches.  Chapter 6 6)  Conclusions and  83  Recommendations  Simplified calculations have shown that the use of air instead of "push steam" for discharging of a typical FSL batch would cause an unwanted rise in the copper level in solution. The increase in the copper content of the solution is estimated at 0.35 g/L in case the compressor would shut down once it reaches the desired discharge pressure of 115 psig. Should the discharge pressure be maintained at 115 psig the incremental copper concentration would be several times higher. Considering the reduced nature of FSL residues, the effect of agitation is trivial; the cuprous sulfide slurry can be expected to react with any oxygen introduced to the reactor almost instantaneously, whether it is stirred or not. Even if the increase in the copper concentration was acceptable, the use of air would be unsafe in view of the possible presence of H S and H in the autoclave freeboard after a 2  copper-depleted leach.  2  84  References and Bibliography  REFERENCES A N DBIBLIOGRAPHY  [1]  I.S. Grewal, "Oxidative Pressure Leaching of Chalcocite in Sulphuric Acid" (M.A.Sc. thesis, The University of British Columbia, October 1991).  [2]  I. Grewal, D.B. Dreisinger, D. Krueger, P.M. Tyroler, E . Krause and N.C. Nissen, "Total Oxidative Leaching of Cu S-containing Residue at INCO Ltd.'s Copper Refinery: 2  Laboratory Studies on the Reaction Pathways," Hvdrometallurgy, 29 (1992), 319-333. [3]  C. Clement, P.M. Tyroler, E . Krause and N.C. Nissen, "Total Oxidative Leaching of Cu S-containing Residue at INCO Ltd.'s Copper Refinery: The Effect of Morphology and 2  Viscosity  of  Basic  Copper  Sulphate  on  the  Leaching  of  Cupric  Sulphide,"  Hvdrometallurgy. 29 (1992), 335-356. [4]  B.R. Saito, "Oxidative Pressure Leaching of Chalcocite by INCO Ltd.'s Second Stage Leach Process" (M.A.Sc. thesis, The University of British Columbia, August 1995).  [5]  B.R. Saito, D.B. Dreisinger, C. Clement, E . Krause and N.C. Nissen, "Total Oxidative Leaching of Cu S-containing Residue at INCO Ltd.'s Copper Refinery: The Effect of 2  Temperature and Feed Arsenic Content on Leaching Behaviour" (Paper to be presented at Copper 95 - Cobre 95, Santiago, Chile, 26-29 November, 1995). [6]  R.J.M. van Lier, "Laboratory Investigations into INCO Ltd.'s C C C R - C R E D First Stage Leach Process" (Progress Report #1, The University of British Columbia, July 1994).  [7]  R.J.M. van Lier, "Laboratory Investigations into INCO Ltd.'s C C C R - C R E D First Stage Leach Process" (Progress Report #2, The University of British Columbia, October 1994).  [8]  J.R. Boldt, Jr., The Winning of Nickel (Toronto, Ont: Longmans Canada Ltd., 1967).  [9]  T. Papademetriou and J.R. Grasso, "Recovery of Precious Metals from South African Matte," Engelhard Industries Technical Bulletin. 10 (4) (1970), 121-129.  [10]  D . M . Chizhikov et al, "Electrolytic Method for the Processing of Copper-Nickel-Cobalt Matte," U.S.S.R. Patent 158.074, 1963, transl. from Russian by J.D. Mcintosh.  85  References and Bibliography [11]  Anon., "Procede pour l'Electroraffinage Direct d'une Anode formee par une Matte de Nickel," French Patent 1.360.675, 1964.  [12]  M . L . Goble and J.A. Chapman, "Electrolytic Nickel Production in the Thompson Refinery," Nickel Metallurgy, Vol. I, E . Ozberk and S.W. Marcuson, eds. (Montreal, Que: CIM, 1986), 464-480.  [13]  Y . Yang and X . Meng,' "Operating Practice and Technical Developments in Nickel Refining  and  Cobalt  Recovery  at  Jinchuan  Non-Ferrous  Metal  Company,"  Electrometallurgical Plant Practice, P.L. Claessens and G.B. Harris, eds. (Montreal, Que: CIM, 1990), 253-268. [14]  T. Inami, Y . Ishikawa, N . Tsuchida and T. Sugiura, "Chlorine Leach of Anode Slime arisen from the Nickel Matte Electrorefining," Extractive Metallurgy of Nickel and Cobalt, G.P. Tyroler and C A . Landolt, eds. (Warrendale, PA: TMS, 1988), 413-427.  [15]  A.R. Burkin, ed., Extractive Metallurgy of Nickel (London, United Kingdom: Society of Chemical Industry, 1987).  [16]  P.G. Thornhill, E . Wigstol and G. Van Weert, "The Falconbridge Matte Leach Process," Journal of Metals. 23 (7) (1971), 13-18.  [17]  F.R. Archibald, "The Kristiansand Nickel Refinery," Journal of Metals, 14 (9) (1962), 648-652.  [18]  L.R. Hougen, R. Parkinson, J. Saetre and G. Van Weert, "Operating Experiences with a Pilot Plant for the Electrowinning of Nickel from All-Chloride Electrolyte," CIM Bulletin, 70 (782) (1977), 136-143.  [19]  E.O. Stensholt, H. Zachariasen, J.H. Lund and P.G. Thornhill, "Recent Improvements in the Falconbridge Nickel Refinery," Extractive Metallurgy of Nickel and Cobalt, G.P. Tyroler and C A . Landolt, eds. (Warrendale, PA: TMS, 1988), 403-412.  [20]  E.O. Stensholt, H . Zachariasen and J.H. Lund, "Falconbridge Chlorine Leach Process," Transactions of The Institution of Mining and Metallurgy, 95 (1986), C10-C16.  86  References and Bibliography [21]  L.R. Hougen and H . Zachariasen, "Recovery of Nickel, Copper and Precious Metal Concentrate from High Grade Precious Metal Mattes," Journal of Metals, 27 (5) (1975), 6-9.  [22]  J.M. Demarfhe, "Elaboration de Nickel Electrolytique a partir de Matte de Nickel," Chloride Hydrometallurgy - Hydrometallurgie des Chlorures (Brussels, Belgium: Benelux Metallurgie, 1977), 231-248.  [23]  Anon., "New Nickel Refinery is back on stream at Le Havre-Sandouville," Engineering and Mining Journal 181 (2) (1980), 35-36.  [24]  R. Crnojevich, D.H. Wilkinson and J.L. Blanco, "Contributions to the Hydrometallurgy of Nickel, Copper and Cobalt in the A M A X , Inc., Port Nickel Plant," Hydrometallurgy: Research, Development and Plant Practice, K. Osseo-Asare and J.D. Miller, eds. (Warrendale, PA: TMS, 1983), 941-953.  [25]  P.M. Tyroler, T.S. Sanmiya and E.W. Hodkin, "Hydrometallurgical Processing of INCO's Pressure Carbonyl Residue," Extractive Metallurgy of Nickel and Cobalt, G.P. Tyroler and C A . Landolt, eds., Warrendale, PA: TMS, 1988), 391-401.  [26]  C. Clement, "Copper Refinery Electrowinning Department Process Description" (revised June, 1993).  [27]  E . Krause (N.C. Nissen),  "Fundamentals  of  the  CRED  Leaching  Operations"  (memorandum to B.R. Conard, February 15, 1989). [28]  H . Saarinen and M . Seilo, "Production of the Cathode Nickel in the Outokumpu Process," Refining Processes in Metallurgy - Raffinationsverfahren  in der  Metallurgie  (Weinheim, Germany: Gesellschaft Deutscher Metallhutten- und Bergleute,  1983),  267-283. [29]  P. Koskinen, M . Virtanen and H. Eerola, "Integrated Nickel Production in Outokumpu Oy," Extractive Metallurgy of Nickel and Cobalt, G.P. Tyroler and C A . Landolt, eds. (Warrendale, PA: TMS, 1988), 355-371.  References and Bibliography [30]  87  S. Fugleberg, S.-E. Hultholm, L . Rosenback and T. Holohan, "Development of the Hartley Platinum Leaching Process" (Paper to be presented at The Principles and Practice of Leaching, Winnipeg, Man., 15-18 October, 1995).  [31]  J.D.G. Groom, R.J.E. Stewart, J.L. Nixon and H. Saarinen, "Development of the Outokumpu Nickel Refining Process in Zimbabwe," Mineral Processing and Extractive Metallurgy, M.J. Jones and P. Gill, eds. (London, United Kingdom: I M M , 1984), 381-395.  [32]  J.K. Copping, "Nickel Concentrate Leaching and Refining at the Kwinana, W.A., Nickel Refinery of Western Mining Corporation Ltd.," Mining and Metallurgical Practices in Australasia, J.T. Woodcock, ed. (Victoria, Australia: AusIMM, 1980), 590-594.  [33]  J.E. Dutrizac and T.T. Chen, "A Mineralogical Study of the Phases Formed during the C u S 0 - H S 0 - 0 Leaching of Nickel-Copper Matte," Canadian Metallurgical Quarterly, 4  2  4  2  26 (4) (1987), 265-276. [34]  R.P. Plasket and S. Romanchuk, "Recovery of Nickel and Copper from High-Grade Matte at Impala Platinum by the Sherritt Process," Hydrometallurgy, 3 (1978), 135-151.  [35]  R.P. Plasket and G . M . Dunn, "Iron Rejection and Impurity Removal from Nickel Leach Liquor at Impala Platinum Limited," Iron Control in Hydrometallurgy, J.E. Dutrizac and A.J. Monhemius, eds. (London, United Kingdom: Society of Chemical Industry, 1986), 695-718.  [36]  Z. Hofirek and P. Halton, "Production of High Quality Electrowon Nickel at Rustenburg Base Metal Refiners (Pty.) Ltd.," Electrometallurgical Plant Practice, P.L. Claessens and G.B. Harris, eds. (Montreal, Que: CIM, 1990), 233-251.  [37]  Z. Hofirek and D.G.E. Kerfoot, "The Chemistry of the Nickel-Copper Matte Leach and its Application to Process Control and Optimisation," Hydrometallurgy, 29 (1992), 357-381.  [38]  Z. Hofirek and P.J. Nofal, "Pressure Leach Capacity Expansion using Oxygen enriched Air at R B M R (Pty.) Ltd." (Paper to be presented at The Principles and Practice of Leaching, Winnipeg, Man., 15-18 October, 1995).  88  References and Bibliography [39]  C.F. Brugman and D . G . Kerfoot, "Treatment of Nickel-Copper Matte at Western Platinum by the Sherritt Acid Leach Process," Nickel Metallurgy, Vol. I, E . Ozberk and S.W. Marcuson, eds. (Montreal, Que: CIM, 1986), 512-531.  [40]  R.M.G.S. Berezowsky, M.J. Collins, D.G.E. Kerfoot and N . Torres, "The Commercial Status of Pressure Leaching Technology," JOM, 43 (2) (1991), 9-15.  [41]  M.D. Faris, M.J. Moloney and O.G. Pauw, "Computer Simulation of the Sherritt NickelCopper Matte Acid Leach Process," Hydrometallurgy. 29 (1992), 261-273.  [42]  H . Gotze  and D. Lowe, "Drucklaugung von Nickelrohstoffen:  Intensivierung  der Nickelmetallurgie,"  Freiberger  Forschungshefte,  ein Beitrag zur B210  (1979),  117-128. [43]  H . Martens, L . Miiller, M . Rudorf and D. Lowe, "Processing of Secondary Nickel Raw Material," Transactions of The Institution of Mining and Metallurgy, 97 (1988), C163-C166.  [44]  P. Fossi, L . Gandon, C. Bozec  and J.M. Demarthe,  "Refining of High-Nickel  Concentrates." CIM Bulletin. 70 (783) (1977), 188-197. [45]  A.J. Bard, R. Parsons and J. Jordan, eds., Standard Potentials in Aqueous Solution (Oxford, United Kingdom: IUPAC, 1985).  [46]  D.D. Wagman et al., Selected Values of Chemical Thermodynamic  Properties,  Technical Note 270-3 (Washington, DC: NBS, 1968). [47]  R.G. Robins, "Arsenic Hydrometallurgy," Arsenic Metallurgy: Fundamentals and Applications, R.G. Reddy, J.L. Hendrix and P.B. Queneau, eds. (Warrendale, PA: T M S , 1988), 215-247.  [48]  M . Pourbaix, ed., Atlas d'Equilibres Electrochimiques  (Atlas of Electrochemical  Equilibria, transl. by J.A. Franklin, Houston, TX: N A C E , 1974). [49]  Y . K . Rao, Stoichiometry and Thermodynamics of Metallurgical Processes (New York, NY: Cambridge University Press, 1985).  89  References and Bibliography [50]  I. Barin, Thermochemical Data of Pure Substances, Part I and JJ (Weinheim, Germany: V C H Verlagsgesellschaft mbH, 1989).  [51]  R.A. Robie, B.S. Hemingway and J.R. Fisher, Thermodynamic Properties of Minerals and Related Substances at 298.15K and 1 Bar (10 Pascals) Pressure and at Higher 5  Temperatures, Geological Survey Bulletin 1452 (Washington, DC: USGS, 1978). [52]  D.J. Vaughan and J.R. Craig, Mineral Chemistry of Metal Sulfides (New York, N Y : Cambridge University Press, 1978).  [53]  C M . Criss and J.W. Cobble, "The Thermodynamic Properties of High Temperature Aqueous Solutions.  IV.  Entropies of the Ions up to 200° and the Correspondence  Principle," Journal of the American Chemical Society, 86 (1964), 5385-5390. [54]  C M . Criss and J.W. Cobble, "The Thermodynamic Properties of High Temperature Aqueous Solutions. V . The Calculation of Ionic Heat Capacities up to 200°. Entropies and Heat Capacities above 200°," Journal of the American Chemical Society, 86 (1964), 5390-5393.  [55]  J.W. Cobble, "The Thermodynamic Properties of High Temperature Aqueous Solutions. VI. Applications of Entropy Correspondence to Thermodynamics and Kinetics," Journal of the American Chemical Society. 86 (1964), 5394-5401.  [56]  W . M . Latimer, The Oxidation States of the Elements and their Potentials in Aqueous Solutions (Englewood Cliffs, NJ: Prentice-Hall, Inc., 1938), 359-369.  [57]  H . C Helgeson, "Thermodynamics of Complex Dissociation in Aqueous Solutions at Elevated Temperatures," Journal of Physical Chemistry, 71 (10) (1967), 3121-3136.  [58]  H . E . Barner and R.V. Scheuerman, Handbook of Thermochemical Compounds and Aqueous Species (New York, NY: John Wiley & Sons, 1978).  [59]  G. Bornemann, "liber die Zusammensetzung  von Scheeles Grim," Zeitschrift fur  Anorganische und Allgemeine Chemie, 124 (1922), 36-39.  90  References and Bibliography [60]  D.A. Huggins, "General Operation of the Copper Refinery Electrowinning Department" (Process Technology Report #5907, February 19, 1973).  [61]  Anon., "Formulating High-Performance Cleaning Products with D O W F A X Anionic Surfactants" (brochure of The Dow Chemical Company).  [62]  K . N . Bech, "Compressed Air for 1 Concerned, November 17, 1992).  st  Stage Autoclave Discharge" (memorandum to A l l  Appendix A  Thermochemical  Calculations  91  APPENDIX A  THERMOCHEMICAL  I  CALCULATIONS  S p e c i e s and thermodynamic data  Tables A - l through A-4 show the complete set of species considered and thermodynamic data used in the calculations of Chapters 3 and 5. Table A - l .  Thermodynamic data of  water  species;  1  on ''conventional"  scale,  § calculated by CSIRO software. o  0  Species  $298K  Source  (calc'd)  f,433K  AG  J/mole/K  kJ/mole  69.950  -237.141  [50]  -§  130.680  0  [50]  0  205.147  0  [50]  0  011  0  [45]  0  -10.75  -157.293  [45]  -119.9  H0 2  o  2  OH-  Table A-2.  f,298K  AG  kJ/mole  Thermodynamic data of sulfur species; 1J before extension metastability,  of sulfur  § destabilized prior to Criss and Cobble calculations,  f classified as a simple anion. 0  Species  $298K  0  f,298K  AG  J/mole/K  kJ/mole  Source  (calc'd)  f,433K  AG  kJ/mole  HS  (g)  205.753  -33.328  [50]  -38.284  HS  (aq)  129  -27.87  [45]  -  62.9  12.05  [45]  30.3 |  -16.3  86.31  [45]  111.4  32.056  0  [50]  0  131.8  -756.01 %  [45]  -397.4 §  20.1  -744.63 U  [45]  -362.5 §  2  2  s2  S (rhombic) HS0  4  so 2  4  Appendix A  92  Thermochemical Calculations  Table A-3.  Thermodynamic data of arsenic species; ^ estimated value, § classified as a simple cation. o  Species  0  Source  $298K  & °f,433K  (calc'd)  G  J/mole/K  kJ/mole  222.782  69.109  [50]  71.101  As (rhombohedral)  35.706  0  [50]  0  AsO HAs0  12.52 Tl  -163.8  [45]  -156.7 §  126.6  -402.7  [45]  -373.1  AsO]  41.2  -350.2  [45]  -304.4  H As0  196.6  -639.9  [45]  -597.0  H AsO}  110.4  -587.5  [45]  -528.2  HAsOf  -35.1H  -524.3  [47]  -435.5  AsO\-  -248.9 If  -447.7  [47]  -323.9  184.1  -766.1  [46]  -707.0  H As0  117.0  -748.5  [45]  -705.0  HAsOf  3.76  -707.1  [45]  -642.1  AsO /  -162.7  -647.5  [45]  -554.4  (claudetite)  117.001  -576.641  [50]  -541.548  (realgar)  254.053  -133.004  [50]  -129.357  -70.32  [51]  -  -166.172  [50]  -164.679  AsH  3  +  2  3  3  2  H As0 3  4  2  4  3  As 0 2  3  As S 4  4  • 63.51  AsS As S 2  3  (orpiment)  Table A-4.  163.594  kJ/mole  Thermodynamic data of nickel species. 0  o  Species  Source  $298K  J/mole/K  kJ/mole  29.874  0  [50]  NiOH  -71.1  -227.6  [46]  Ni  2+  -159.0  -46.4  [45]  HMO]  -  -349.2  [48]  NiO (bunsenite)  37.991  -211.539  [50]  Ni S  133.888  -210.396  [50]  53.011  -85.205  [50]  Ni +  3  2  (heazlewoodite)  NiS (millerite)  Appendix A II  Thermochemical  93  Calculations  Entropy estimations  Arsenyl ion: Latimer's method According to Latimer [56], the standard entropy of a complex aqueous species can be approximated by adding the individual entropy contributions of the element and the ligand. Whereas the former is usually known, the latter may be calculated from available data for similar species. Thus, for the arsenyl ion, the entropy contribution of the ligand (oxygen) is calculated from the antimonyl ion: Table A-5.  Thermodynamic data used in the calculation of the standard entropy of the arsenyl ion. 0  complex aqueous  S  0  298K  (species)  S  298K  (element)  (Ugand)  °298K  S  J/mole/K  J/mole/K  J/mole/K  SbO  +  22.33  45.522  -23.19  AsO  +  ?  35.706  -23.19  species  Hence: S (AsO ) +  298K  = 35.706 + (-23.19)  = 12. 52 J. mole' .K' 1  1  (A-l)  If more entropy data for similar species are available they all should be used to calculate a more accurate average value for the entropy contribution of the ligand.  Di-ortho-arsenite and ortho-arsenite ions It is assumed that the stepwise decrease in entropy for the ortho-arsenious acid dissociation series is proportionally the same as for arsenic acid:  Appendix A Table A-6.  Thermochemical  94  Calculations  Thermodynamic data used in the calculation of the standard entropy of the di-ortho-arsenite and ortho-arsenite ions.  Species  0  0  ^298K  $298K  Species  J/mole/K  J/mole/K  H As0  184.1  196.6  H As0  H As0  117.0  110.4  H As0  3.76  ?  HAsOf  -162.7  ?  AsOf  3  4  2  4  HAs0  2  As0  4  3 4  3  2  3  3  Hence: S° o (HAs0 -)  = 110.4  2  7()  K  298KK  S°  298K  III  —— 184.1-117.0  3S  (AsO\-)  = -35.1  - '™ 'j' * S  (196.6-110.4)  1  1  (A-2) ' y  ^ ( %- 6-110.4)  16  ]  4  = - 35.1 J. mole' .K'  = - 248.9 J. mole' .K' 1  1  (A-3)  C r i s s and Cobble's theory  Correct application of the Criss and Cobble theory requires the formation reactions of all charged aqueous species from their elements be written as full-cell reactions. In this way ambiguities in defining the properties of electrons are avoided. As an example, consider the half-cell reaction for the formation of the arsenyl ion: As + H 0  -» AsO  +  2  + 2H  +  + 2e"  (A-4)  Equation (A-4) is combined with the hydrogen half-cell reaction: H  + e  +  -> JH  (A-5)  2  and the water formation reaction: H  2  to give:  +  j  0  2  -> H 0 2  (A-6)  Appendix A  Thermochemical Calculations  95_  As + j 0 + H -» AsO + j H +  (A-7)  +  2  2  Since, by definition, the Gibbs free energies of formation of the elements and protons are zero at 25°C, it follows that: *Gf,298K  AG°  =  (AsO ) +  (A-8)  f>298K  The formation reactions for the other ions are written in a similar fashion:  As + 0 + j H 2  As0 + H  (A-9)  +  2  2  As + | 0 + J H -» H AsO} + H  +  2  2  (A-10)  2  •As + j 0 + j H -> HAsOf + 2H  +  2  2  As + j 0 + j H 2  ^5 +  +  0 + j H -> H AsO' 2  As + 0  2  2  + | H  2  ^  AsOf + 3 H  2  3  S° + H  ^  2  HST + H  S - + 2H 2  2  5 ° + 2 <9 + tf 2  2  + H  + H  2  ^  (A-13) (A-14)  (A-15)  (A-17)  +  -> flSO; + #  +  (A-18)  -> SOf + 2H  (A-19)  +  2  (A-12)  (A-16)  +  S° + H  2  +  +  2  2  j 0  +  H -> As0 i + 3 H  + <9 + j  2  + //  -> /£4jO^ + 2 / /  2  5° + 2 0  4  (A-ll)  OH' + H  +  (A-20)  Appendix A  Thermochemical  Calculations  96  Criss and Cobble's entropy correspondence principle [53] is mathematically expressed as: S°  = a(T)  TK  + b(T)S  (A-21)  298K  The tabulated coefficients a and b are functions of both temperature and the type of ion [53,54]. Equation (A-21) is only satisfied when the "conventional" standard entropies of the ions (in cal/mole/K) are converted to the "absolute" scale [54] through the relationship:  S°298K( )  =  abs  "  °298K  S  5  -  0  (A-22)  z  wherein z is the ionic charge. The average heat capacity of ions is given by:  \TK  _ S°  -  TK  298K  S° (abs) 298K  ,  (A-23)  T  ln-  298.15  The great usefulness of this extension of the entropy correspondence principle, of course, is that it does not require any direct information on the heat capacity for electrolytes at any temperature. The average heat capacities of the non-ionic species - except for undissociated complexes - are o  calculated from existing C data: p  TK  \c (T) p  T  K =  298K  dT  298K  (  T-298.15  v  A  _  2  4  )  '  wherein: C° (T)  = A + BT + CT  (A-25)  2  The heat capacity functions (in J/mole/K) used in the spreadsheet calculations are taken from Rao [49]: C (T,H ) 2  = 27.280 + 3.264xlQ- T+ 3  0.502xl0 r 5  2  (A-26)  Appendix A  Thermochemical C° (T,0 ) p  Calculations  = 29.957  2  C (T,As)  p  + 4.184 x 10' T - 1.674.x. 10 T 3  s  2  = 23.179 + 5.523xl()- T  (A-27)  (A-28)  3  p  C (T,S)  97_  = 14.811 + 24.058 x 10~ T + 0.728 x l 0 T 3  5  2  (A-29)  The change in heat capacity of the ion formation reaction is obtained by summing the average o  heat capacities of products minus reactants.  AGy - for the ion formation reaction can then rA  readily be calculated according to: TK  AS  298K  (T - 298.15)  + ACJ  ©  (A-30)  with:  0  = T - 298.15 - Tin— 298.15  (A-31)  Since the free energies of the elements and protons at all temperatures are zero by convention, o  AGy  T K  for the reaction is equal to the Gibbs free energy of formation of the ion at temperature o  T. The calculated values of AGj-  433K  IV  for all ions considered are reported in Tables A - l to A-3.  Helgeson's theory  In Helgeson's analysis [57] the total interaction between complex, ions and solvent dipoles contributing to the entropy change of dissociation of an undissociated species in aqueous solution is regarded as the sum of electrostatic (long range) and non-electrostatic (short range) interaction: AS (diss) T  = AS° (T) e  + AS° (T) n  = AS (T )f(T) e  r  + AS° (T )g(T) n  r  (A-32)  wherein T is the reference temperature (25°C). r  The combination of a general form of the Born or Bjerrum equation and a mathematical relationship expressing the variation of the dielectric constant of water with temperature in the 0-370°C interval gives the following equation for the electrostatic entropy term:  Appendix A  Thermochemical  AS (T)  AS (T )exp exp(b + aT) - oxp(b + aT ) +  =  e  98  Calculations  e  0  r  r  1 + a©exp(b  ^ (A-33)  + aT)  1 + a@Qxp(b + aT ) r  with a = 0.01875, b = -12.741 and 0 = 219. AS° (T) is equal to the net difference in the entropy change attending solvation of the neutral e  complex and its dissociated species as long as the hydration process is entirely electrostatic. The standard state adopted in Helgeson's theory is based on infinite dilution. Consequently, ion-ion o  and ion-complex interaction are negligible, and AS (T)  refers solely to the electrostatic  e  interaction of water dipoles with the complex and its dissociated ions. Assuming that for dissociational reactions in aqueous solutions the non-electrostatic contribution to the heat capacity of dissociation can be represented by:  = a + PT+ r  AC„ (T)  (A-34)  Y  n  in which a, B and y are reaction-dependent coefficients, then equation (A-34) can be divided by T and integrated to give the following approximation to the non-electrostatic entropy term:  ASJT)  = AS (T ) n  •+ c d n ^  r  + B(T-T ) r  +  (A-35)  -(T -T?)  J  2  Substitution of equations (A-33) and (A-35) into equation (A-32) yields:  AS (diss) T  =  AS (T )exp exp(b + aT) - txp(b + aT ) + e  r  r  1 + a@exp(b + aT)  + AS (T ) n  1 + a© exp(b + aT )  r  + aln^  T-T r  0  (A-36)  + B(T-T ) r  +  -(T -T )  Y  2  2  r  r  The general expression for the dissociation constant of dissociational reactions as a function of temperature resulting from this analysis is cumbersome and unpractical. However,  \ogK (diss) T  for many complexes below «200°C can be closely approximated by assuming that: ASJT)  ASJT)  AS (T ) e  r  AS  n  (T ) r  (A-37)  Appendix A  Thermochemical  99  Calculations  which is consistent with the statement that: o  A C j (diss) AC° JT)  (A-38)  constant  p  "  p  The effectiveness of this approximation is partly due to the insensitivity of \ogK  (diss) to  T  substantial departures from the above constant ratio, and partly to the similar behavior of the non-electrostatic power function of temperature and the electrostatic exponential function at low temperatures. The assumption that A C  log K (diss) =  o p T  (T) leads to:  r  ^  CO  r  ±  M (T ) T  pe  - — 1 - exp exp(b + aT) - cxp(b + aT ) +  T  T  o  (diss) is proportional to AC  r  (A-39) AH (T )  +  T  RTlnlO  r  RTlnlO  where: co = 1 + aQexp(a  + bT ) r  = 1.00322  (A-40)  Only when the entropy and enthalpy of dissociation at 25°C are both negative, close approximations of logK  (diss) (or AG (diss))  T  AC° (diss)lAC° (T) pT  T  to about 200°C can be made assuming  is constant. The negative entropy value reflects the major contribution  pe  of solvent interaction to complex stability. On the contrary, if the heat capacity of dissociation o  o  o  and (or) AS (T ) and AH (T ) are positive, the non-electrostatic contribution to AC' T  r  T  r  o  r  (diss) is  ,  considerable and an approximation to logK (diss) (or AG (diss)) using equation (A-39) is not T  warranted.  T  In the borderline case, i.e. negative entropy and large positive enthalpy of  dissociation (e.g. the water equilibrium), the approximation of equation (A-38) also holds. In such case the effectiveness of equation (A-39) is primarily due to the fact that logK (diss) dominated by.the enthalpy term. T  is  Since: AGj(diss)  = -RTlnK (diss) T  = -RTlnlOlogK  T  (diss)  (A-41)  Appendix A  Thermochemical  Calculations  100  it follows that:  AG (diss) T  =  T  - — 1 - exp exp(b + aT) - exp(b + aT ) +  AS (T ) T  r  CO  r  r  T-T  +  AH (T ) T  0  -  (A-42)  r  Hence, the free energy of formation of the neutral complex is calculated using the equation: o  AGfj  ^  (complex)  o  o  = 2_,AGfj(ions)  - AGj(diss)  (A-43)  The free energies of formation of the ions are calculated by the method of Criss and Cobble as previously described. The reactions treated according to the simplified Helgeson equations are: HAsQ  -> H  +  2  H AsQ  3  -> H  H As0  4  -> H  3  3  +  +  (A-44)  + AsO'  2  +  H AsO'  +  H As0  2  3  2  4  (A-45) (A-46)  Not considered are the dissociation reactions: H (g)  ->  H (aq)  (AAT)  0 (g)  ->  0 (aq)  (A-48)  ^  H S(aq)  (A-49)  ->  AsH (aq)  (A-50)  2  2  H S(g) 2  AsH (g) 3  2  2  2  3  For these reactions involving molecular gases the situation is more complicated due to large nonelectrostatic contributions. Moreover, the_ CSIRO Thermochemistry software utilizes data for non-hydrated oxygen and hydrogen gas for the computation of the water stability lines at all temperatures.  Appendix B  Experiment Design Calculations  101  APPENDIX B  EXPERIMENT DESIGN  CALCULATIONS  CRED PLANT DATA Densities: IPC slurry IPC solids water electrolyte  1.89E+03 5.5E+03 1.0E+03 1.26E+03  kg/m3 {dilution due to pump flange v\ kg/m3 kg/m3 kg/m3  Constant Density Tank: 6550 5255 5255 1295 7125 57.6%  L L kg = L kg = (wt.)  600 600 12500 15750 33.0%  L kg = L kg = (wt.)  {heating to 80°C} 1323 lbs {mixture of spent and pregnar 34722 lbs {at the end of preheating}  steam condensate volume steam mass total batch mass total batch volume  3500 3500 32230 23150  L kg = kg = L  {heating to 160°C} 7716 lbs 71053 lbs {at the beginning of leaching}  % solids in autoclave % water in autoclave % electrolyte in autoclave  5.6%» (vol.) 22.1% (wt.) 29.0% (wt.) 40.4% (vol.) 48.9% (wt.) + 54.0% (vol.) + 100% 100%  tank volume water volume water mass solids volume solids mass % solids in slurry  11584 lbs 15708 lbs  First-Stage Batch Make-Up Tank: steam condensate volume steam mass electrolyte volume electrolyte mass % solids in batch First-Stage  Autoclave:  LABORATORY BATCH MAKE-UP MAWL add solids add water add electrolyte  1.3 0.073 0.525 0.702  L L= L= L=  {from Parr manuals} 400 g take 360 g 525 mL take 500 mL 702 mL take 700 mL  not included}  Appendix C  102  Experiment Worksheets APPENDIX C  EXPERIMENT  WORKSHEETS  This appendix contains the worksheets of autoclave runs and releaching experiments carried out as part of the present investigations, in chronological order. The results of both unsuccessful tests and experiments which are now considered irrelevant are not included. In case mechanical difficulties were encountered during a certain test, the nature of the problem is specified on the worksheet.  There was generally not enough sample available to repeat those experiments.  It  goes without saying that data generated by such tests were utilized with prudence in Chapter 5. The reported assays are a collection of chiefly ICP analyses, copper electrodeposition tests, acid titrations and L E C O analyses, the vast majority of which was performed by INCO's Copper Cliff Central Process Technology Laboratory. A small number of solutions and residues generated during INCO's summer shutdown were analyzed at the International Plasma Laboratory (IPL) in Vancouver.  Appendix C  Experiment Worksheets  103  RELEACHING TEST WORKSHEET  test ID: date:  #F Monday, July 10,1995  medium:  NaOH  natural pH: 6.40 testpH: 12.00 temperature: leaching time at T:  50 ° C 60 min  initial pulp density:  10% solids  122 ° F 1h  REAGENT BALANCE Beaker feed  Beaker products  FSL residue #62: deionized water: 1M NaOH solution:  33.3 g 300 mL 53.6 mL  leach residue: leach filtrate: wash water: total filtrate: repulp volume:  ARSENIC ASSAYS time minutes 0 30 60  leach sol'n ppm 0 347 396 456  solids repulp sol'n ppm ppm 28833  25068  ARSENIC BALANCE time minutes  -  0 30 60  As in beaker As in sol'n As in solids mg mg mg 960.14 0 960.14 960.14 122.68 837.46 956.67 136.05 820.62 952.71 152.01 800.70  Assay head As extraction  12.4%  As distribution solution solids 0.0% 100.0% 12.8% 87.2% 14.2% 85.8% 16.0% 84.0%  33.6 298 205 503  g mL mL mL  438 mL  Appendix C  104  Experiment Worksheets  RELEACHING TEST WORKSHEET  test ID:  #E  date:  Monday, July 10,1995  medium: NaOH natural pH: 6.75 testpH: 11.00 temperature: leaching time at T:  50 ° C 60 min  initial pulp density:  10% solids  122 ° F 1h  REAGENT BALANCE Beaker products  Beaker feed FSL residue #62: deionized water: 1M NaOH solution:  33.3 g 300 mL 7.8 mL  leach residue: leach filtrate: wash water: total filtrate: repulp volume:  ARSENIC ASSAYS time minutes 0 30 60  leach sol'n ppm 0 0 277 367  solids repulp sol'n ppm ppm 28833 27255  0  ARSENIC BALANCE time minutes 0 30 60  As in beaker As in sol'n As in solids mg mg mg 960.14 0 960.14 960.14 0 960.14 960.14 82.49 877.65 957.37 105.56 851.81  Assay head As extraction  6.8%  As distribution solution solids 0.0% 100.0% 0.0% 100.0% 8.6% 91.4% 11.0% 89.0%  32.9 255 215 470  g mL mL mL  401 mL  Appendix C  105  Experiment Worksheets  RELEACHING TEST WORKSHEET  test ID:  #D  date:  Friday, July 7,1995  medium: NaOH natural pH: 4.18 testpH: 11.98 temperature: leaching time at T:  50 ° C 60 min  initial pulp density:  2 0 % solids  REAGENT  122 ° F 1 h  BALANCE Beaker products  Beaker feed FSL residue #63: deionized water: 1M NaOH solution:  75 g 300 mL 61.1 mL  leach residue: leach filtrate: wash water: total filtrate:  74.8 g 283 mL 232 mL 515 mL  repulp volume:  437 mL  ARSENIC A S S A Y S time minutes 0 30 60  leach sol'n ppm 0 14 21 21.8  solids repulp sol'n ppm ppm 5298 4969  0  ARSENIC B A L A N C E time minutes 0 30 60  As in beaker As in sol'n As in solids mg mg mg 397.35 0 397.35 5.05 397.35 392.30 397.21 7.37 389.84 397.00 7.45 389.55  Assay head As extraction  6.5%  As distribution solution solids 0.0% 100.0% 98.7% 1.3% 1.9% 98.1% 1.9% 98.1%  Appendix C  106  Experiment Worksheets  RELEACHING TEST WORKSHEET  test ID:  #C  date:  Friday, July 7,1995  medium: NaOH natural pH: 3.77 testpH: 11.01 temperature: leaching time at T:  50 ° C 60 min  initial pulp density:  2 0 % solids  122 ° F 1h  REAGENT BALANCE Beaker products  Beaker feed FSL residue #63: deionized water: 1M NaOH solution:  75 g 300 mL 10.7 mL  leach residue: leach filtrate: wash water: total filtrate: repulp volume:  ARSENIC ASSAYS time minutes -  0 30 60  leach sol'n ppm 0 0 0 0  repulp sol'n solids ppm ppm 5298 5753  0  ARSENIC BALANCE time minutes 0 30 60  As in beaker As in sol'n As in solids mg mg mg 397.35 0 397.35 397.35 0 397.35 397.35 0 397.35 397.35 0 397.35  Assay head As extraction  -6.0%  As distribution solution solids 0.0% 100.0% 0.0% 100.0% 0.0% 100.0% 0.0% 100.0%  73.2 235 140 375  g mL mL mL  333 mL  107  Appendix C Experiment Worksheets COMBINED AUTOCLAVE AND RELEACHING TEST WORKSHEET date:  Friday, June 30, 1995  AUTOCLAVE TEST test ID: series ID:  RELEACHING TEST  #65 Releaching  temperature: heat-up time: leaching time:  160 ° C 49 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar 13.7  stirring rate:  test ID:  #B  medium:  "Durco filtrate" with  temperature:  320 1 I  leaching time:  100 g / L C u S 0 4 130 g/L H 2 S 0 4 176 ° F  80 ° C 180 min  3 h  85 psig 85 psig 820 rpm  1/s  REAGENT BALANCE Test products  Test feed  releach residue: releach filtrate: wash solution: total filtrate:  IPC residue: electrolyte: deionized water:  360 g 700 m L 500 m L  "Durco filtrate":  150 m L a t 6 5 ° C repulp volume:  274 957 421 1378  g mL mL mL  498 m L  SLURRY PbtEWTlALS before autoclave leaching: after autoclave leaching: releaching:  65 mV (SCE)  310 mV (SHE)  -168 mV (SCE)  77 m V (SHE) at 8 0 ° C  85 mV (SCE)  330 mV (SHE) at 8 0 ° C  ASSAYS IPC residue May 21,199 g/L 55.21% 35 7.05% 8.75% 0.79% 10 5.09% 6.82% 15.15%  electrolyte component Cu Co Ni As Fe O Stot H2S04  releach residue 74.86% 0.70% 0.55% 2.79% 0.26%  time minutes 0 5 15 30 60 120 180  17.74%  220  ARSENIC BALANCE time minutes  0 5 15 30 60 120 180  A s in liner g 8.013 8.013 8.012 8.010 8.007 8.005 8.003 8.000  A s in sol'n g 0.030 0.230 0.273 0.282 0.296 0.320 0.369 0.498  As in solids g 7.984 7.784 7.738 7.728 7.712 7.686 7.634 7.502  A s distribution solution solids 0.4% 99.6% 2.9% 97.1% 3.4% 96.6% 3.5% 96.5% 3.7% 96.3% 4.0% 96.0% 4.6% 95.4% 6.2%  93.8%  releach solution ppm A s 22 170 204 212 224 244 284 386  repulp filtrate ppm A s  8  Appendix C  108  Experiment Worksheets  COMBINED A U T O C L A V E A N D R E L E A C H I N G T E S T W O R K S H E E T  date:  Thursday, June 29,1995 RELEACHING T E S T  AUTOCLAVE TEST #64 Releaching  test ID: series ID:  160 ° C 49 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  REAGENT  #A  medium:  "Durco filtrate" with  temperature:  320  temperature: heat-up time: leaching time:  stirring rate:  test ID:  leaching time:  1  80 ° C 180 min  3 h  BALANCE Test products  Test feed  releach residue: releach filtrate: wash solution: total filtrate:  IPC residue: electrolyte: deionized water:  360 g 700 mL 500 mL  "Durco filtrate":  220 mL at 6 5 ° C  286 1083 347 1430  g mL mL mL  542 mL  repulp volume: SLURRY  100 g / L C u S 0 4 130 g/L H 2 S 0 4 176 ° F  P6TENTIALS  before autoclave leaching: after autoclave leaching: releaching:  65 mV (SCE)  310 mV (SHE)  -310 mV (SCE)  -65 mV (SHE) at 8 0 ° C  85 mV (SCE)  330 mV (SHE) at 8 0 ° C  ASSAYS  IPC residue May 21,199 g/L 55.21% 35 7.05% 8.75% 0.79% 10 5.09% 6.82% 15.15% 220  electrolyte component Cu Co Ni As Fe O Stot H2S04 ARSENIC time minutes  0 5 15 30 60 120 180  releach residue 74.35% 0.93% 0.79% 3.55% 0.33%  time minutes  0 5 15 30 60 120 180  17.74%  BALANCE As in liner g 10.271 10.271 10.270 10.270 10.270 10.269 10.269 10.268  As in sol'n As in solids g 0.007 0.048 0.044 0.050 0.060 0.073 0.096 0.146  g 10.264 10.222 10.227 10.220 10.210 10.196 10.173 10.122  As distribution solution solids 0.1% 99.9% 0.5% 99.5% 0.4% 99.6% 0.5% 99.5% 0.6% 99.4% 0.7% 99.3% 0.9% 99.1% 1.4% 98.6%  releach solution ppm A s 5 34 31 36 43 53 70 107  repulp filtrate ppm A s  6  Appendix C  109  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #61 Arsenic Saturday, May 13,1995  temperature: heat-up time to T: leaching time at T:  160 ° C 50 min 240 min  initial pressure: final pressure:  5.8 bar 5.8 bar 13.7  stirring rate: REAGENT  1/s  320 ° F 4 h 84 psig 84 psig 820 rpm  BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  360 g 700 mL 500 mL  leach residue:  336 g  residue ratio:  leach filtrate: wash solution: total filtrate:  930 mL 1014 mL 1944 mL  wash acidity:  repulp volume:  746 mL  1.07 wt./wt.  11 g / L H 2 S 0 4  P6TENTIAL M £ A S U R £ M £ N T S before:  slurry:  after:  slurry: supernatant:  60 mV (SCE)  305 mV (SHE)  -332 mV (SCE) -186 mV (SCE)  -87 mV (SHE) 59 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O  electrolyte 35 g/L  10 g/L  Stot H2S04  component Cu Co Ni As Fe H2S04  leach filtrate 0 ' 28.5 25.9 0 22.2  g/L g/L g/L g/L g/L  57.3 g/L IPC residue of May 24, 1994 10.28 3.49 2.94  0 0.484 0.442 0 0.398  M M M M M  0.584 M  repulp filtrate 0 0.137 0.289 0.004 0.108  g/L g/L g/L g/L g/L  11 g/L  leach residue 69.48 4.18 16.63  BALANCE assay head extraction  consumption kg/t IPC residue 68.1  99.2% 74.1% 98.9% 248.9  detailed arsenic balance 9.664  As in As out difference  leach residue 73.65% 0.06% 2.19% 3.21% 0.05% 1.06% 17.63%  220 g/L  element ratio (wt./wt.) Cu : 0 C u : Stot Stot: O METALLURGICAL  IPC residue of May 24,1994 57.04% 7.10% 7.90% 0.74% 4.03% 5.55% 16.34%  g 10.805 g 1.141 g  A s precipitated A s in leach r e s . : A s in IPC res.  10.56% 100% 4.05 (wt./wt.)  calc'd head grade 61.93% 7.43% 8.75% 1.06% 5.79%  calc'd head extraction 99.2% 76.6% 99.2%  balance errors (w.r.t. head grades) absolute relative 4.89% 8.58% 0.33% 4.66% 0.85% 10.72% 0.32% 42.83% 1.76% 43.68%  Appendix C  110  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET test ID: series ID: date:  #60 Arsenic Thursday, May 11,1995  temperature: heat-up time to T: leaching time at T:  160 ° C 51 min 60 min  initial pressure: final pressure:  5.8 bar 5.8 bar 13.7  stirring rate:  1/s  320 ° F 1 h 84 psig 84 psig 820 rpm  REAGENT BALANCE Autoclave feed  Autoclave products 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  P6TENTIAL  leach residue:  310 g  residue ratio:  leach filtrate: wash solution: total filtrate:  949 mL 971 mL 1920 mL  wash acidity:  repulp volume:  704 mL  1.16 wt./wt.  11 g / L H 2 S 0 4  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  60 mV (SCE)  305 mV (SHE)  -305 mV (SCE) -86 mV (SCE)  -60 mV (SHE) 159 mV (SHE)  ASSAYS component Cu Co Ni As Fe O  electrolyte 35 g/L  10 g/L  Stot H2S04  IPC residue of May 24,1994 57.04% 7.10% 7.90% 0.74% 4.03% 5.55% 16.34%  leach residue 74.87% 0.50% 1.43% 3.17% 0.19% 1.05% 18.39%  220 g/L  g/L g/L g/L g/L g/L  61.8 g/L IPC residue of May 24,1994 10.28 3.49 2.94  element ratio (wt./wt.) Cu : 0 C u : Stot Stot: O  leach filtrate 0.05 24.8 27.9 0.01 18.7  0.001 0.421 0.475 0.000 0.335  M M M M M  0.630 M  repulp filtrate 0.072 0.171 0.202 0.008 0.125  g/L g/L g/L g/L g/L  11 g/L  leach residue 71.30 4.07 17.51  METALLURGICAL BALANCE component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 67.9  93.9% 84.4% 95.9% 235.2  detailed arsenic balance  As in A s out difference  9.664 g 9.826 g 0.162 g  A s precipitated As in leach r e s . : A s in IPC res.  1.65% 99.9% 3.69 (wt./wt.)  calc'd head grade 57.68% 6.97% 8.58% 0.79% 5.10%  calc'd head extraction 93.8% 85.6% 96.7%  balance errors (w.r.t. head grades) absolute relative 0.64% 1.12% 0.13% 1.79% 0.68% 8.58% 0.05% 6.10% 1.07% 26.50%  Appendix C  111  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET test ID: series ID: date:  #59 Arsenic Wednesday, May 10,1995  temperature: heat-up time to T: leaching time at T:  160 ° C 49 min 30 min  initial pressure: final pressure:  5.8 bar 5.8 bar  stirring rate:  320 ° F 0.5 h 84 psig 84 psig 820 rpm  13.7 1/s  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  360 g 700 mL 500 mL  leach residue:  308 g  . residue ratio:  leach filtrate: wash solution: total filtrate:  927 mL 1020 mL 1947 mL  wash acidity:  repulp volume:  627 mL  1.17 wt./wt.  11 g / L H 2 S 0 4  POTENTIAL MEASUREMENTS before:  slurry:  after:  slurry: supernatant:  60 mV (SCE)  305 mV (SHE)  -75 mV (SCE) 177 mV (SCE)  170 mV (SHE) 422 mV (SHE)  ASSAYS component Cu Co Ni As Fe O Stot H2SQ4  electrolyte 35 g/L  10 g/L  IPC residue of May 24,1994 57.04% 7.10% 7.90% 0.74% 4.03% 5.55% 16.34%  leach residue 75.81% 1.11% 0.58% 2.81% 0.39% 0.97% 18.63%  leach filtrate 0.01 24.7 30.0 0.66 17.5  220 g/L  66.5 g/L IPC residue of May 24,1994 10.28 3.49 2.94  element ratio (wt./wt.) Cu : 0 C u : Stot Stot: O  g/L g/L g/L g/L g/L  0.000 0.419 0.511 0.009 0.313  M M M M M  0.678'M  repulp filtrate 0.009 0.451 0.552 0.092 0.351  g/L g/L g/L g/L g/L  12 g/L  leach residue 78.15 4.07 19.21  METALLURGICAL BALANCE component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 68.0  86.6% 93.7%  calc'd head grade 58.06% 7.30% 8.22%  91.6%  0.63% 4.83% 225.3  detailed arsenic balance As in As out difference  9.664 g 9.267 g 0.397 g  A s precipitated As in leach r e s . : As in IPC res.  4.29% 91.3% 3.25 (wt./wt.)  .  calc'd head extraction 87.0% 93.9% 93.0%  balance errors (w.r.t. head grades) absolute relative 1.02% 1.78% 0.20% 2.84% 0.32% 4.05% 14.92% 0.11% 0.80% 19.92%  Appendix C  Experiment Worksheets  112  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #58 Arsenic Thursday, May 5,1995  temperature: heat-up time to T: leaching time at T:  160 ° C 51 min 20 min  initial pressure: final pressure:  5.8 bar 5.8 bar  stirring rate: REAGENT  320 ° F 0.33 h 84 psig 84 psig  13.7 1/s  820 rpm  BALANCE  Autoclave feed  Autoclave products  IPC residue: electrolyte: deionized water:  360 g 700 mL 500 mL  leach residue:  307 g  residue ratio:  leach filtrate: wash solution: total filtrate:  965 mL 945 mL 1910 mL  wash acidity:  repulp volume:  617 mL  1.17 wt./wt.  11 g / L H 2 S 0 4  POTENTIAL M E A S U R E M E N T S before:  slurry:  after:  slurry: supernatant:  60 mV (SCE)  305 mV (SHE)  -41 m V ( S C E ) 135 mV (SCE)  204 mV (SHE) 380 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 35 g/L  10 g/L  component Cu Co Ni As Fe H2S04  leach filtrate 0.17 22.1 29.2 4.92 15.9  g/L g/L g/L g/L g/L  72.6 g/L IPC residue of May 24, 1994 10.28 3.49 2.94  0.003 0.374 0.497 0.066 0.285  M M M M M  0.740 M  repulp filtrate 0.191 0.130 0.142 0.062 0.105  g/L g/L g/L g/L g/L  11 g/L  leach residue 53.94 4.07 13.25  BALANCE assay head extraction  consumption kg/t IPC residue 67.6  81.2% 91.4% 89.2% 204.4  detailed arsenic balance A s in As out difference  leach residue 75.51% 1.56% 0.80% 1.63% 0.51% 1.40% 18.55%  220 g/L  element ratio (wt./wt.) Cu : 0 C u : Stot Stot : O METALLURGICAL  IPC residue of May 24,1994 57.04% 7.10% 7.90% 0.74% 4.03% 5.55% 16.34%  9.664 g 9.747 g 0.083 g  As precipitated As in leach r e s . : A s in IPC res.  0.85% 32.1% 1.87 (wt./wt.)  calc'd head grade 57.63% 7.25% 8.50% 0.76% 4.70%  calc'd head extraction 81.6% 92.0% 90.7%  balance errors (w.r.t. head grades) absolute relative 0.59% 1.04% 0.15% 2.04% 0.60% 7.64% 0.02% 3.12% 0.67% 16.58%  Appendix C  Experiment Worksheets  113  AUTOCLAVE TEST WORKSHEET test ID: series ID: date:  #57 Arsenic Thursday, May 4, 1995  temperature: heat-up time to T: leaching time at T:  160 ° C 51 min 10 min  initial pressure: final pressure:  5.8 bar 5.8 bar  320 ° F 0.17 h 84 psig 84 psig  13.7 1/s  stirring rate:  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  360 g 700 mL 500 mL  leach residue:  315 g  residue ratio:  leach filtrate: wash solution: total filtrate:  956 mL 899 mL 1855 mL  wash acidity:  repulp volume:  719 mL  1.14 wt./wt.  11 g / L H 2 S Q 4  POTENTIAL MEASUREMENTS before:  slurry:  after:  slurry: supernatant:  60 mV (SCE)  305 mV (SHE)  -40 mV (SCE) 127 mV (SCE)  205 mV (SHE) 372 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2SQ4  electrolyte 35 g/L  10 g/L  component Cu Co Ni As Fe H2S04  leach filtrate 0.26 21.2 34.1 4.26 15.7  g/L g/L g/L g/L g/L  69.4 g/L IPC residue of May 24,1994 10.28 3.49 2.94  0.004 0.360 0.581 0.057 0.281  M M M M M  0.708 M  repulp filtrate 0.282 0.300 0.453 0.128 0.250  g/L g/L g/L g/L g/L  11 g/L  leach residue 46.31 4.12 11.25  BALANCE assay head extraction  consumption kg/t IPC residue 67.4  72.8% 87.5%  calc'd head grade 59.32% 7.56% 10.04%  83.8%  0.80% 4.82% 215.9  detailed arsenic balance A s in A s out difference  leach residue 75.49% 2.21% 1.13% 1.84% 0.74% 1.63% 18.33%  220 g/L  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O METALLURGICAL  IPC residue of May 24,1994 57.04% 7.10% 7.90% 0.74% 4.03% 5.55% 16.34%  9.664 g 9.880 g 0.216 g  A s precipitated A s in leach r e s . : A s in IPC res.  2.19% 41.8% 2.18 (wt./wt.)  calc'd head extraction 74.5% 90.2% 86.5%  balance errors (w.r.t. head grades) absolute relative 2.28% 3.99% 0.46% 6.50% 2.14% 27.10% 0.06% 8.11% 0.79% 19.70%  Appendix C  114  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET test ID: series ID: date:  #56 Arsenic Wednesday, May 3,1995  temperature: heat-up time to T: leaching time at T:  160 ° C 50 min 0 min  initial pressure: final pressure:  5.8 bar 5.8 bar  320 ° F 0 h 84 psig 84 psig  13.7 1/s  stirring rate:  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  360 g 700 mL 500 mL  leach residue:  319 g  residue ratio:  leach filtrate: wash solution: total filtrate:  972 mL 869 mL 1841 mL  wash acidity:  repulp volume:  738 mL  1.13 wt./wt.  11 g / L H 2 S 0 4  PdTENTIAL MEASUREMENTS before:  slurry:  60 mV (SCE)  305 mV (SHE)  after:  slurry: supernatant:  32 mV (SCE) 42 mV (SCE)  277 mV (SHE) 287 mV (SHE)  ASSAYS component Cu Co Ni As Fe O Stot H2S04  electrolyte 35 g/L  10 g/L  IPC residue of May 24,1994 57.04% 7.10% 7.90% 0.74% 4.03% 5.55% 16.34%  leach residue 72.67% 3.48% 1.98% 0.82% 1.09% 2.32% 17.84%  220 g/L  g/L g/L g/L g/L g/L  81.2 g/L IPC residue of May 24,1994 10.28 3.49 2.94  element ratio (wt./wt.) Cu : 0 C u : Stot Stot: O  leach filtrate 2.96 14.7 32.6 7.19 13.5  0.047 0.249 0.556 0.096 0.241  M M M M M  0.828 M  repulp filtrate 0.350 0.036 0.041 0.013 0.031  g/L g/L g/L g/L g/L  11 g/L  leach residue 31.32 . 4.07 7.69  METALLURGICAL BALANCE component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 60.1  56.6% 77.8%  10.56% 0.72% 4.60%  75.9% 181.9  detailed arsenic balance As in As out difference  9.664 g 9.606 g 0.058 g  A s precipitated As in leach res. : A s in IPC res.  calc'd head grade 58.39% 7.04%  0.60% 0.19% 0.98 (wt./wt.)  calc'd head extraction 56.2% 83.4% 78.9%  balance errors (w.r.t. head grades) absolute relative 1.35% 2.36% 0.06% 0.84% 2.66% 33.62% 0.02% 2.18% 0.57% 14.24%  Appendix C  AUTOCLAVE TEST  test ID:  115  Experiment Worksheets WORKSHEET  #55  series ID: date:  Additives Saturday, April 29,  1995  160 °C  temperature:  320 ° F  heat-up time to T :  50 min  leaching time at T :  60 min  1 h  initial pressure:  5.9 bar  85 psig  final pressure:  5.9 bar  85 psig  13.7  stirring rate:  820 rpm  1/s  REAGENT BALANCE A u t o c l a v e products  A u t o c l a v e feed IPC residue:  360 g  electrolyte:  800 m L  d e i o n i z e d water:  400 m L  leach residue:  308 g  leach filtrate:  959 m L  w a s h solution:  871 m L  w a s h acidity:  1.17 wt./wt. •  11 g / L H 2 S 0 4  1830 m L  total filtrate: repulp volume:  POTENTIAL  residue ratio:  696 m L  MEASUREMENTS 67mV(SCE)  before:  slurry:  after:  slurry: supernatant:  312 m V ( S H E )  263 m V ( S C E )  -18 m V ( S H E )  89 m V ( S C E )  334 m V ( S H E )  ASSAYS IPC residue of component  electrolyte  May 10,  1994  leach  leach  residue  filtrate  repulp filtrate  54.90%  77.45%  .0.06 g/L  0.001  Co  7.07%  0.58%  26.8 g/L  0.456 M  0.181  Ni  9.76%  0.67%  37.4 g/L  0.638 M  0.296 g/L  As  0.73%  0.82%  0 g/L  0 M  0 g/L  Fe  4.46%  0.25%  18.2 g/L  0.326 M  0.125 g / L  95.2 g / L  0.970 M  11 g / L  45 g / L  Cu  O  6.75%  1.62%  Stot  14.83%  18.26%  220 g / L  H2S04  M  0.079 g/L g/L  5 ppm  Crodaglu  100 p p m  Dowfax  element ratio (wt./wt.) Cu:0  IPC residue of  leach  May 10,1994  residue  8.13  47.81  C u : Stot  3.70  4.24  Stot:0  2.20  11.27  METALLURGICAL BALANCE  component  assay head  consumption  calc'd h e a d  calc'd h e a d  b a l a n c e errors (w.r.t. h e a d grades)  extraction  kg/t IPC residue  grade  extraction  absolute  Cu  99.8  56.28%  1.38%  relative 2.51%  Co  92.9%  7.65%  93.5%  0.58%  8.22%  Ni  94.1%  10.55%  94.5%  0.79%  8.06%  As  3.8%  0.70%  0.0%  0.03%  3.77%  Fe  95.3%  5.06%  95.8%  0.60%  13.56%  H2S04  208.7  Appendix C  116  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #54 Additives Friday, April 28, 1995 160 °C 51 min 60 min  temperature: heat-up time to T: leaching time at T: initial pressure: final pressure: stirring rate:  320 °F 1 h  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 800 mL 400 mL  IPC residue: electrolyte: deionized water:  P6TENTIAL  leach residue:  310 g  residue ratio:  leach filtrate: wash solution: total filtrate:  948 mL 882 mL 1830 mL  wash acidity:  repulp volume:  664 mL  1.16 wt./wt.  11 g/L H2S04  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  66 mV (SCE)  311 mV (SHE)  -259 mV (SCE) 76 mV (SCE)  -14mV(SHE) 321 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04 Crodaglu Dowfax  electrolyte 45 g/L  IPC residue of May 10,1994 54.90% 7.07% 9.76% 0.73% 4.46% 6.75% 14.83%  leach residue 77.32% 0.55% 0.94% 0.89% 0.25% 1.50% 18.4%  220 g/L 0 ppm 0 ppm IPC residue of May 10,1994 8.13 3.70 2.20  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot : O  leach filtrate 0.06 g/L 25.6 g/L 36.1 g/L Og/L 17.0 g/L  0.001 M 0.434 M 0.615 M 0M 0.305 M  94.3 g/L  0.962 M  repulp filtrate 0 0.112 0.269 0 0.077  g/L g/L g/L g/L g/L  11 g/L  leach residue 51.55 4.20 12.27  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 99.8  93.3% 91.7% -4.6% 95.2% 213.6  calc'd head grade 56.60% 7.21% 10.31% 0.76% 4.70%  calc'd head extraction 93.4% 92.2% 0.0% 95.4%  balance errors (w.r.t. head grades) absolute relative 1.70% 3.09% 0.14% 1.92% 0.55% 5.67% 0.03% 4.64% 0.24% 5.46%  Appendix C  AUTOCLAVE TEST  test ID: series ID: date:  117  Experiment Worksheets WORKSHEET  #53 Kinetics, Low Cu : S Thursday, April 27, 1995  temperature: heat-up time to T: leaching time at T:  160 °C 51 min 15 min  initial pressure: final pressure:  5.9 bar 5.9 bar  stirring rate:  13.7 1/s  320 °F 0.25 h 85 psig 85 psig 820 rpm  REAGENT BALANCE Autoclave feed  Autoclave products 360 g 800 mL 400 mL  IPC residue: electrolyte: deionized water:  POTENTIAL  leach residue:  306 g  residue ratio:  leach filtrate: wash solution: total filtrate:  980 mL 785 mL 1765 mL  wash acidity:  repulp volume:  738 mL  1.18 wt./wt.  11 g/L H2S04  MEASUREMENTS  before:  slurry:  69mV(SCE)  314 mV (SHE)  after:  slurry: supernatant:  50 mV (SCE) 55 mV (SCE)  295 mV (SHE) 300 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 19,1994 53.59% 7.52% 8.93% 0.77% 5.39% 5.88% 16.73%  leach residue 73.31% 2.38% 1.18% 0.47% 0.80% 1.06% 20.12%  220 g/L  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  IPC residue of May 19, 1994 9.11 3.20 2.85  leach filtrate 11.2 20.6 31.1 1.44 17.9  g/L g/L g/L g/L g/L  0.176 M 0.350 M 0.530 M 0.019 M 0.321 M  88.7 g/L  0.904 M  repulp filtrate 0.251 0.144 0.188 0.013 0.122  g/L g/L g/L g/L g/L  11 g/L  leach residue 69.16 3.64 18.98  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 69.6  73.1% 88.8% 47.6% 87.4% 223.5  calc'd head grade 55.36% 7.64% 9.47% 0.79% 5.56%  calc'd head extraction 73.5% 89.4% 49.2% 87.8%  balance errors (w.r.t. head grades) absolute relative 1.77% 3.29% 0.12% 1.62% 0.54% 6.06% 0.02% 3.20% 0.17% 3.08%  Appendix C  118  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #52 Kinetics, Low Cu : S Wednesday, April 26,1995 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 50 min 30 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  stirring rate:  0.5 h  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 800 mL 400 mL  IPC residue: electrolyte: deionized water:  P6TENTIAL  leach residue:  314 g  residue ratio:  leach filtrate: wash solution: total filtrate:  977 mL 829 mL 1806 mL  wash acidity:  repulp volume:  641 mL  1.15 wt./wt.  11 g/L H2S04  MEASUREMENTS  before:  slurry:  69 mV (SCE)  314 mV (SHE)  after:  slurry: supernatant:  34 mV (SCE) 42 mV (SCE)  279 mV (SHE) 287 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 19, 1994 53.59% 7.52% 8.93% 0.77% 5.39% 5.88% 16.73%  leach residue 74.29% 1.41% 0.78% 0.38% 0.64% 0.94% 20.12%  220 g/L IPC residue of May 19, 1994 9.11 3.20 2.85  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate 4.2 24.8 33.7 1.80 20.6  g/L g/L g/L g/L g/L  0.066 M 0.421 M 0.574 M 0.024 M 0.368 M  92.4 g/L  0.942 M  repulp filtrate 0.125 0.063 0.081 0.008 0.057  g/L g/L g/L g/L g/L  11 g/L  leach residue 79.03 3.69 21.40  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 88.6  83.7% 92.4% 56.8% 89.6% 212.8  calc'd head grade 55.94% 7.96% 9.82% 0.82% 6.14%  calc'd head extraction 84.6% 93.1% 59.5% 90.9%  balance errors (w.r.t. head grades) absolute relative 2.35% 4.38% 0.44% 5.87% 0.89% 10.00% 0.05% 6.61% 0.75% 13.87%  Appendix C  Experiment Worksheets  AUTOCLAVE TEST  test ID: series ID: date:  119  WORKSHEET  #51 Kinetics, Low Cu : S Tuesday, April 25, 1995 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 50 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave feed  Autoclave products  IPC residue: electrolyte: deionized water:  POTENTIAL  360 g 800 mL 400 mL  leach residue:  310 g  residue ratio:  leach filtrate: wash solution: total filtrate:  981 mL 941 mL 1922 mL  wash acidity:  repulp volume:  711 mL  1.16 wt./wt.  11 g/L H2S04  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  70 mV (SCE)  315 mV (SHE)  -9 mV (SCE) 143 mV (SCE)  236 mV (SHE) 388 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 19,1994 53.59% 7.52% 8.93% 0.77% 5.39% 5.88% 16.73%  leach residue 72.48% 0.79% 0.37% 0.45% 0.27% 0.78% 19.61%  220 g/L IPC residue of May 19, 1994 9.11 3.20 2.85  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate 0.09 25.7 33.3 1.21 22.0  g/L g/L g/L g/L g/L  0.001 M 0.436 M 0.568 M 0.016 M 0.393 M  89.3 g/L  0.911 M  repulp filtrate 0.006 0.022 0.032 0.077 0.015  g/L g/L g/L g/L g/L  11 g/L  leach residue 92.92 3.70 25.14  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 99.7  91.0% 96.4% 50.0% 95.7% 216.7  calc'd head grade 52.44% 7.68% 9.41% 0.72% 6.21%  calc'd head extraction 91.2% 96.6% 46.2% 96.3%  balance errors (w.r.t. head grades) absolute relative 1.15% 2.15% 0.16% 2.07% 0.48% 5.32% 0.05% 7.14% 0.82% 15.30%  Appendix C  Experiment Worksheets  AUTOCLAVE TEST  test ID: series ID: date:  120  WORKSHEET  #50 Kinetics, Low Cu : S Thursday, April 20, 1995 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 49 min 0 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  stirring rate:  0 h  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 800 mL 400 mL  IPC residue: electrolyte: deionized water:  leach residue:  317 g  residue ratio:  leach filtrate: wash solution: total filtrate:  926 mL 812 mL 1738 mL  wash acidity:  repulp volume:  727 mL  before:  slurry:  68mV(SCE)  313 mV (SHE)  after:  slurry: supernatant:  62 mV (SCE) 48 mV (SCE)  307 mV (SHE) 293 mV (SHE)  1.14 wt./wt.  11 g/L H2S04  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 19, 1994 53.59% 7.52% 8.93% 0.77% 5.39% 5.88% 16.73%  leach residue 67.31% 3.86% 1.87% 0.66% 1.23% 1.69% 18.97%  220 g/L IPC residue of May 19,1994 9.11 3.20 2.85  element ratio (wtVwt.) Cu : 0 Cu : Stot Stot : O  leach filtrate 21.9 18.3 31.5 1.04 17.7  g/L g/L g/L g/L g/L  0.344 M 0.311 M 0.536 M 0.014 M 0.318 M  90.4 g/L  0.922 M  repulp filtrate 0.361 0.121 0.178 0.009 0.136  g/L g/L g/L g/L g/L  11 g/L  leach residue 39.83 3.55 11.22  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 43.8  54.8% 81.6% 24.6% 79.9% 231.5  calc'd head grade 54.90% 8.11% 9.74% 0.85% 5.65%  calc'd head extraction 58.1% 83.1% 31.5% 80.8%  balance errors (w.r.t. head grades) absolute relative 1.31% 2.44% 0.59% 7.83% 0.81% 9.12% 0.08% 10.01% 0.26% 4.78%  Appendix C AUTOCLAVE TEST  test ID: series ID: date:  121  Experiment Worksheets WORKSHEET  #49 Kinetics, Low Cu : S Wednesday, April 19, 1995 320 °F  160 °C 51 min 225 min  temperature: heat-up time to T: leaching time at T: initial pressure: final pressure: stirring rate:  loose agitator 3.75 h  5.5 bar 5.5 bar  80 psig 80 psig  13.7 1/s  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  POTENTIAL  360 g 800 mL 400 mL  leach residue:  321 g  residue ratio:  leach filtrate: wash solution: total filtrate:  954 mL 902 mL 1856 mL  wash acidity:  repulp volume:  629 mL  1.12 wt./wt.  11 g/L H2S04  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  79 mV (SCE)  324 mV (SHE)  16mV(SCE) 138 mV (SCE)  261 mV (SHE) 383 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 19,1994 53.59% 7.52% 8.93% 0.77% 5.39% 5.88% 16.73%  leach residue 73.85% 0.80% 0.41% 0.77% 0.24% 0.75% 19.79%  220 g/L IPC residue of May 19,1994 9.11 3.20 2.85  element ratio (wt./wt.) Cu:0 Cu : Stot Stot: O  leach filtrate Og/L 27.3 g/L 35.7 g/L 0.16 g/L 23.4 g/L  0M 0.463 M 0.608 M 0.002 M 0.419 M  83.0 g/L  0.846 M  repulp filtrate 0 0.093 0.131 0.075 0.077  g/L g/L g/L g/L g/L  11 g/L  leach residue 98.47 3.73 26.39  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 100.0  90.5% 95.9% 10.5% 96.0% 241.4  calc'd head grade 55.85% 7.95% 9.82% 0.73% 6.41%  calc'd head extraction 91.0% 96.3% 5.8% 96.7%  balance errors (w.r.t. head grades) absolute relative 2.26% 4.22% 0.43% 5.68% 0.89% 9.98% 0.04% 4.96% 1.02% 19.00%  Appendix C  122  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #48 Additives Wednesday, April 12,1995 160 °C 550 min 60 min  temperature: heat-up time to T: leaching time at T: initial pressure: final pressure: stirring rate:  320 °F 1h  5.7 bar 6.6 bar  82 psig 96 psig  13.7 1/s  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  P6TENTIAL  leach residue:  319 g  residue ratio:  leach filtrate: wash solution: total filtrate:  946 mL 961 mL 1907 mL  wash acidity:  repulp volume:  797 mL  1.13 wt./wt.  11 g/L H2S04  MEASUREMENTS 63 mV (SCE)  before:  slurry:  after:  slurry: supernatant:  -257 mV (SCE) 163 mV (SCE)  308 mV (SHE) -12 mV (SHE) 408 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04 Crodaglu Dowfax  electrolyte 45 g/L  IPC residue of May 10,1994 54.90% 7.07% 9.76% 0.73% 4.46% 6.75% 14.83%  leach residue 74.83% 0.57% 1.52% 0.90% 0.25% 2.31% 17.84%  220 g/L 5 ppm 100 ppm IPC residue of May 10, 1994 8.13 3.70 2.20  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate Og/L 29.2 g/L 37.2 g/L Og/L 19.3 g/L  0M 0.495 M 0.634 M 0 M 0.345 M  71.6 g/L  0.730 M  repulp filtrate 0.001 0.300 0.736 0 0.196  g/L g/L g/L g/L g/L  11 g/L  leach residue 32.39 4.19 7.72  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 87.5  92.9% 86.2% -9.2% 95.0% 210.1  calc'd head grade 57.56% 8.18% 11.13% 0.80% 5.29%  calc'd head extraction 93.8% 87.9% 0.0% 95.8%  balance errors (w.r.t. head grades) absolute relative 2.66% 4.84% 1.11% 15.71% 1.37% 14.04% 0.07% 9.25% 0.83% 18.62%  Appendix C  Experiment Worksheets  AUTOCLAVE TEST  test ID: series ID: date:  123  WORKSHEET  #47 Additives Tuesday, April 11, 1995 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 55 min 60 min  initial pressure: final pressure:  5.7 bar 6.6 bar  82 psig 95 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  POTENTIAL  360 g 700 mL 500 mL  leach residue:  314 g  residue ratio:  leach filtrate: wash solution: total filtrate:  945 mL 945 mL 1890 mL  wash acidity:  repulp volume:  646 mL  1.15 wt/wt.  11 g/L H2S04  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  64 mV (SCE)  309 mV (SHE)  -246 mV (SCE) 133 mV (SCE)  -1 mV (SHE) 378 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04 Crodaglu Dowfax  electrolyte 45 g/L  IPC residue of May 10, 1994 54.90% 7.07% 9.76% 0.73% 4.46% 6.75% 14.83%  leach residue 75.55% 0.58% 1.44% 0.91% 0.22% 1.78% 17.93%  220 g/L 0 ppm 0 ppm IPC residue of May 10, 1994 8.13 3.70 2.20  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate Og/L 28.2 g/L 32.8 g/L Og/L 19.1 g/L  0 M 0.479 M 0.558 M 0M 0.342 M  69.5 g/L  0.709 M  repulp filtrate 0.001 0.648 1.110 0 0.451  g/L g/L g/L g/L g/L  11.5 g/L  leach residue 42.44 4.21 10.07  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 87.5  92.8% 87.1% -8.7% 95.7% 216.4  calc'd head grade 57.15% 7.91% 9.86% 0.79% 5.20%  calc'd head extraction 93.6% 87.3% -0.0% 96.3%  balance errors (w.r.t. head grades) absolute relative 2.25% 4.09% 0.84% 11.94% 0.10% 0.98% 0.06% 8.73% 0.74% 16.67%  Appendix C  Experiment Worksheets  AUTOCLAVE TEST  test I D : series I D : date:  124  WORKSHEET  #46 Kinetics, High Cu : S Friday, April 7, 1995 80 °C 31 min 240 min  temperature: heat-up time to T: leaching time at T:  4 h  bar bar  psig psig  13.7 1/s  820 rpm  initial pressure: final pressure: stirring rate:  176 °F  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  POTENTIAL  360 g 700 mL 500 mL  leach residue:  335 g  residue ratio:  leach filtrate: wash solution: total filtrate:  949 mL 901 mL 1850 mL  wash acidity:  repulp volume:  758 mL  1.07 wt./wt.  11 g / L H 2 S 0 4  MEASUREMENTS  before:  slurry:  62 mV (SCE)  307 mV (SHE)  after:  slurry: supernatant:  65 mV (SCE) 68 mV (SCE)  310 mV (SHE) 313 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2SQ4  electrolyte 45 g/L  IPC residue of May 11, 1994 57.91% 7.40% 8.89% 0.72% 4.73% 6.28% 13.32%  leach residue 66.03% 6.59% 4.72% 0.63% 2.60% 3.03% 15.52%  220 g/L  IPC residue of May 11,1994 9.22 4.35 2.12  element ratio (wt./wt.) Cu:0 C u : Stot Stot: O  leach filtrate 31.0 7.5 20.1 0.71 12.4  g/L g/L g/L g/L g/L  0.489 M 0.128 M 0.342 M 0.010 M 0.222 M  84.8 g/L  0.865 M  repulp filtrate 0.80 0.09 0.24 0.011 0.13  g/L g/L g/L g/L g/L  11.5 g/L  leach residue 21.79 4.25 5.12  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 5.7  17.2% 50.6% 18.7% 48.8% 176.7  calc'd head grade 60.88% 8.12% 9.68% 0.77% 5.70%  calc'd head extraction 24.5% 54.7% 24.4% 57.5%  balance errors (w.r.t. head grades) absolute relative 2.97% 5.13% 0.72% 9.68% 0.79% 8.90% 0.05% 7.43% 0.97% 20.46%  Appendix C  Experiment Worksheets  125  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #45 Kinetics, High Cu : S Thursday, April 6, 1995  temperature: heat-up time to T: leaching time at T:  80 °C 31 min 0 min  initial pressure: final pressure:  176 °F 0 h  bar bar  stirring rate:  13.7 1/s  psig psig 820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  PdTENTIAL  leach residue:  354 g  residue ratio:  leach filtrate: wash solution: total filtrate:  958 mL 895 mL 1853 mL  wash acidity:  repulp volume:  752 mL  1.02 wt./wt.  11 g/L H2SQ4  MEASUREMENTS  before:  slurry:  62 mV (SCE)  307 mV (SHE)  after:  slurry: supernatant:  68 mV (SCE) 73 mV (SCE)  313 mV (SHE) 318 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O  electrolyte 45 g/L  Stot H2S04  IPC residue of May 11, 1994 57.91% 7.40% 8.89% 0.72% 4.73% 6.28% 13.32%  leach residue 60.35% 7.30% 8.27% 0.62% 4.25% 4.40% 14.36%  220 g/L  IPC residue of May 11,1994 9-22 4.35 2.12  element ratio (wt./wt.) Cu:0 Cu : Stot Stot: O  leach filtrate 35.9 2.9 5.3 0.57 4.8  g/L g/L g/L g/L g/L  0.565 M 0.049 M 0.091 M 0.008 M 0.085 M  124.4 g/L  1.268 M  repulp filtrate 1.11 0.06 0.13 0.011 0.12  g/L g/L g/L g/L g/L  12 g/L  leach residue 13.72 4.20 3.26  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kgA IPC residue -8.0  3.0% 8.5% 16.0% 11.6% 69.5  calc'd head grade 60.14% 7.95% 9.55% 0.76% 5.45%  calc'd head extraction 9.6% 14.8% 20.1% 23.2%  balance errors (w.r.t. head grades) absolute relative 2.23% 3.85% 0.55% 7.38% 0.66% 7.46% 0.04% 5.19% 0.72% 15.14%  Appendix C  126  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #44 Kinetics, High Cu : S Wednesday, April 5, 1995  temperature: heat-up time to T: leaching time at T:  80 °C 30 min 60 min  initial pressure: final pressure:  176 °F 1h  bar bar  stirring rate:  13.7 1/s  psig psig 820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  P6TENTIAL  360 g 700 mL 500 mL  leach residue:  331 g  residue ratio:  leach filtrate: wash solution: total filtrate:  954 mL 841 mL 1795 mL  wash acidity:  repulp volume:  762 mL  1.09 wt./wt.  11 g/L H2S04  MEASUREMENTS  bofore:  slurry:  62 mV (SCE)  307 mV (SHE)  after:  slurry: supernatant:  64 mV (SCE) 70 mV (SCE)  309 mV (SHE) 315 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 11,1994 57.91% 7.40% 8.89% 0.72% 4.73% 6.28% 13.32%  leach residue 63.61% 6.93% 6.17% 0.63% 3.58% 3.37% 14.96%  220 g/L  IPC residue of May 11,1994 9.22 4.35 2.12  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate 32.0 5.8 14.2 0.61 9.9  g/L g/L g/L g/L g/L  0.504 M 0.099 M 0.243 M 0.008 M 0.177 M  96.5 g/L  0.984 M  repulp filtrate 0.77 0.01 0.16 0.008 0.11  g/L g/L g/L g/L g/L  11.5 g/L  leach residue 18.88 4.25 4.44  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 2.6  13.9% 36.2% 19.7% 30.4% 146.4  calc'd head grade 58.23% 7.92% 9.44% 0.74% 5.91%  calc'd head extraction 19.6% 40.0% 22.0% 44.3%  balance errors (w.r.t. head grades) absolute relative 0.32% 0.54% 0.52% 7.00% 0.55% 6.22% 0.02% 2.93% 1.18% 24.91%  Appendix C  Experiment Worksheets  127  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #43 Kinetics, High Cu : S Tuesday, April 4,1995 80 °C 28 min 480 min  temperature: heat-up time to T: leaching time at T:  8 h  bar bar  psig psig  13.7 1/s  820 rpm  initial pressure: final pressure: stirring rate:  176 °F  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  P6TENTIAL  360 g 700 mL 500 mL  leach residue:  320 g  residue ratio:  leach filtrate: wash solution: total filtrate:  953 mL 916 mL 1869 mL  wash acidity:  repulp volume:  803 mL  1.13 wt./wt.  11 g/L H2S04  MEASUREMENTS  before:  slurry:  63 mV (SCE)  308 mV (SHE)  after:  slurry: supernatant:  63 mV (SCE) 66 mV (SCE)  308 mV (SHE) 311 mV(SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 11,1994 57.91% 7.40% 8.89% 0.72% 4.73% 6.28% 13.32%  leach residue 67.01% 6.37% 4.38% 0.62% 3.91% 3.14% 15.56%  220 g/L  IPC residue of May 11,1994 9.22 4.35 2.12  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 29.8 9.0 22.2 0.73 12.7  g/L g/L g/L g/L g/L  0.470 M 0.153 M 0.379 M 0.010 M 0.227 M  79.4 g/L  0.810 M  leach residue 21.34 4.31 4.96  repulp filtrate 0.51 0.03 0.07 0.004 0.04  g/L g/L g/L g/L g/L  10.5 g/L  •  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 8.5  23.4% 56.2% 23.1% 26.6% 189.6  calc'd head grade 58.72% 8.05% 9.77% 0.75% 6.83%  calc'd head extraction 29.6% 60.2% 25.8% 49.1%  balance errors (w.r.t. head grades) absolute relative 0.81% 1.39% 0.65% 8.77% 0.88% 9.95% 0.03% 3.63% 2.10% 44.35%  Appendix C  Experiment Worksheets  AUTOCLAVE TEST  test ID: series ID: date:  128  WORKSHEET  #42 Kinetics, High Cu : S Monday, April 3, 1995 120 °C 40 min 0 min  temperature: heat-up time to T: leaching time at T:  0 h  1.9 bar 1.9 bar  28 psig 28 psig  13.7 1/s  820 rpm  initial pressure: final pressure: stirring rate:  248 °F  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach residue:  329 g  residue ratio:  leach filtrate: wash solution: total filtrate:  926 mL 868 mL 1794 mL  wash acidity:  repulp volume:  691 mL  before:  slurry:  62 mV (SCE)  307 mV (SHE)  after:  slurry: supernatant:  68 mV (SCE) 74 mV (SCE)  313 mV (SHE) 319 mV (SHE)  1.09 wt./wt.  11 g/L H2S04  ASSAYS  component Cu Co Ni As Fe O  electrolyte 45 g/L  Stot H2S04  IPC residue of May 14,1994 55.10% 7.35% 8.72% 0.71% 4.99% 7.00% 13.42%  leach residue 62.27% 6.70% 5.78% 0.62% 3.97% 4.18% 15.64%  220 g/L  IPC residue of May 14,1994 7.87 4.11 1.92  element ratio (wt./wt.) Cu:0 Cu:Stot Stot:0  leach filtrate 37.2 7.7 17.6 0.72 12.7  g/L g/L g/L g/L g/L  0.585 M 0.131 M 0.300 M 0.010 M 0.227 M  84.2 g/L  0.858 M  repulp filtrate 0.76 0.10 0.22 0.012 0.15  g/L g/L g/L g/L g/L  11 g/L  leach residue 14.90 3.98 3.74  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue -8.1  16.6% 39.4% 19.6% 27.3% 184.7  calc'd head grade 57.72% 8.11% 9.81% 0.76% 6.89%  calc'd head extraction 24.5% 46.1% 24.6% 47.3%  balance errors (w.r.t. head grades) absolute relative 2.62% 4.76% 0.76% 10.38% 1.09% 12.54% 0.05% 6.66% 1.90% 38.11%  Appendix C  Experiment Worksheets  129  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #41 Kinetics, High Cu : S Friday, March 31,1995  temperature: heat-up time to T: leaching time at T:  120 °C 39 min 15 min  initial pressure: final pressure:  2.0 bar 2.0 bar  stirring rate:  13.7 1/s  248 °F 0.25 h 29 psig 29 psig 820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach residue:  312 g  residue ratio:  leach filtrate: wash solution: total filtrate:  956 mL 886 mL 1842 mL  wash acidity:  repulp volume:  784 mL  1.15 wt./wt.  11 g/L H2S04  P6TENTIAL M E A S U R E M E N T S before:  slurry:  64 mV (SCE)  309 mV (SHE)  after:  slurry: supernatant:  63 mV (SCE) 72 mV (SCE)  308 mV (SHE) 317 mV (SHE)  ASSAYS  component Cu Co Ni As Fe  electrolyte 45 g/L  O Stot H2S04  IPC residue of May 14, 1994 55.10% 7.35% 8.72% 0.71% 4.99% 7.00% 13.42%  leach residue 65.68% 5.94% 3.79% 0.63% 2.43% 3.46% 16.09%  220 g/L  IPC residue of May 14,1994 7.87 4.11 1-92  element ratio (wt./wt.) Cu:0 Cu:Stot StOt:0  leach filtrate 30.1 10.6 22.3 0.73 14.4  g/L g/L g/L g/L g/L  0.473 M 0.180 M 0.380 M 0.010 M 0.259 M  75.0 g/L  0.765 M  repulp filtrate 0.71 0.09 0.21 0.009 0.12  g/L g/L g/L g/L g/L  11 g/L  leach residue 18.98 4.08 4.65  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 7.7  30.0% 62.3% 22.6% 57.9% 201.5  calc'd head grade 56.15% 7.96% 9.22% 0.74% 5.94%  calc'd head extraction 35.3% 64.3% 26.2% 64.6%  balance errors (w.r.t. head grades) absolute relative 1.05% 1.91% 0.61% 8.28% 0.50% 5.69% 0.03% 4.81% 0.95% 18.95%  Appendix C  AUTOCLAVE TEST  test ID: series ID: date:  130  Experiment Worksheets WORKSHEET  #40 Kinetics, High Cu : S Thursday, March 30,1995 248 °F  temperature: heat-up time to T: leaching time at T:  120 °C 38 min 30 min  initial pressure: final pressure:  2.0 bar 2.0 bar  29 psig 29 psig  13.7 1/s  820 rpm  stirring rate:  0.5 h  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach residue:  314 g  residue ratio:  1.15 wt./wt.  leach filtrate: wash solution: total filtrate:  953 mL 849 mL 1802 mL  wash acidity:  11.5 g/L H2SQ4  repulp volume:  702 mL  before:  slurry:  63 mV (SCE)  308 mV (SHE)  after:  slurry: supernatant:  66 mV (SCE) 74 mV (SCE)  311 mV(SHE) 319 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 14,1994 55.10% 7.35% 8.72% 0.71% 4.99% 7.00% 13.42%  leach residue 67.29% 5.60% 3.26% 0.62% 2.54% 3.79% 16.60%  220 g/L  IPC residue of May 14,1994 7.87 4.11 1.92  element ratio (wt./wt.) Cu:0 Cu:Stot Stot:0  leach filtrate 30.2 12.5 26.3 0.79 15.4  g/L g/L g/L g/L g/L  0.476 M 0.213 M 0.448 M 0.011 M 0.276 M  69.3 g/L  0.706 M  repulp filtrate 0.72 0.15 0.31 0.012 0.17  g/L g/L g/L g/L g/L  11 g/L  leach residue 17.75 4.05 4.38  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 7.5  33.5% 67.4% 23.2% 55.7% 217.2  calc'd head grade 57.94% 8.20% 9.81% 0.76% 6.30%  calc'd head extraction 40.4% 71.0% 27.8% 64.9%  balance errors (w.r.t. head grades) absolute relative 2.84% 5.16% 0.85% 11.61% 1.09% 12.48% 0.05% 6.36% 1.31% 26.18%  Appendix C  131  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #39 Kinetics, High Cu : S Wednesday, March 29,1995 248 °F  temperature: heat-up time to T: leaching time at T:  120 °C 39 min 60 min  initial pressure: final pressure:  2.0 bar 2.0 bar  29 psig 29 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  360 g 700 mL 500 mL  leach residue:  303 g  residue ratio:  1.19 wt./wt.  leach filtrate: wash solution: total filtrate:  955 mL 845 mL 1800 mL  wash acidity:  11.5 g/L H2S04  repulp volume:  650 mL  P6TENTIAL M E A S U R E M E N T S before:  slurry:  61 mV (SCE)  306 mV (SHE)  after:  slurry: supernatant:  60 mV (SCE) 65 mV (SCE)  305 mV (SHE) 310 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 14, 1994 55.10% 7.35% 8.72% 0.71% 4.99% 7.00% 13.42%  leach residue 67.66% 4.99% 2.56% 0.60% 2.40% 3.39% 16.60%  220 g/L  IPC residue of May 14,1994 7.87 4.11 1.92  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 25.7 14.5 26.6 0.83 15.9  g/L g/L g/L g/L g/L  0.404 M 0.245 M 0.453 M 0.011 M 0.285 M  71.4 g/L  0.728 M  repulp filtrate 0.59 0.15 0.27 0.012 0.13  g/L g/L g/L g/L g/L  11.5 g/L  leach residue 19.96 4.08 4.90  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 19.4  42.8% 75.3% 28.3% 59.5% 211.5  calc'd head grade 55.00% 8.04% 9.21% 0.73% 6.24%  calc'd head extraction 47.7% 76.6% 30.3% 67.6%  balance errors (w.r.t. head grades) absolute relative 0.10% 0.18% 0.69% 9.38% 0.49% 5.58% 0.02% 2.86% 1.25% 25.13%  Appendix C AUTOCLAVE  test ID: series ID: date:  Experiment Worksheets  TEST  132  WORKSHEET  #38 Kinetics, High Cu : S Tuesday, March 28, 1995 120 °C 50 min 360 min  temperature: heat-up time to T: leaching time at T:  REAGENT  leaking cooling valve 6 h  1.9 bar 1.9 bar  28 psig 28 psig  13.7 1/s  820 rpm  initial pressure: final pressure: stirring rate:  248 °F  BALANCE  Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  P6TENTIAL  360 g 700 mL 500 mL  leach residue:  309 g  residue ratio:  1.17 wt./wt.  leach filtrate: wash solution: total filtrate:  953 mL 859 mL 1812 mL  wash acidity:  11.5 g/L H2S04  repulp volume:  634 mL  MEASUREMENTS  before:  slurry:  63 mV (SCE)  308 mV (SHE)  after:  slurry: supernatant:  51 mV(SCE) 54 mV (SCE)  296 mV (SHE) 299 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  component Cu Co Ni As Fe H2S04  leach residue 73.96% 2.07% 1.24% 0.48% 1.55% 2.08% 16.74%  220 g/L  IPC residue of May 14,1994 7.87 4.11 1-92  element ratio (wt./wt.) Cu: O Cu:Stot Stot: O  METALLURGICAL  IPC residue of May 14, 1994 55.10% 7.35% 8.72% 0.71% 4.99% 7.00% 13.42%  leach filtrate 8.2 24.6 31.7 1.24 18.7  g/L g/L g/L g/L g/L  0.129 M 0.417 M 0.540 M 0.017 M 0.334 M  65.0 g/L  0.663 M  repulp filtrate 0.40 0.37 0.46 0.019 0.27  g/L g/L g/L g/L g/L  11.5 g/L  leach residue 35.56 4.42 8.05  BALANCE  assay head extraction  consumption kg/t IPC residue 65.8  75.8% 87.8% 42.3% 73.3% 228.2  calc'd head grade 56.90% 8.28% 9.45% 0.74% 6.27%  calc'd head extraction 78.5% 88.7% 44.5% 78.8%  balance errors (w.r.t. head grades) absolute relative 1.80% 3.27% 0.93% 12.68% 0.73% 8.41% 0.03% 3.86% 1.28% 25.69%  Appendix C  Experiment Worksheets  133  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #37 Kinetics, High Cu : S Friday, February 3, 1995 284 °F  temperature: heat-up time to T: leaching time at T:  140 °C 46 min 15 min  initial pressure: final pressure:  3.6 bar 3.6 bar  52 psig 52 psig  13.7 1/s  820 rpm  stirring rate:  0.25 h  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  POTENTIAL  leach residue:  309 g  residue ratio:  leach filtrate: wash solution: total filtrate:  947 mL 688 mL 1635 mL  wash acidity:  repulp volume:  485 mL  1.17 wt./wt.  12 g/L H2S04  MEASUREMENTS  before:  slurry:  65 mV (SCE)  310 mV (SHE)  after:  slurry: supernatant:  60 mV (SCE) 63 mV (SCE)  305 mV (SHE) 308 mV (SHE)  ASSAYS  component Cu Co Ni As Fe 0 Stot H2S04  electrolyte 45 g/L  IPC residue of May 13,1994 56.32% 7.49% 8.79% 0.68% 5.01% 7.20% 12.50%  leach residue 68.0% 4.40% 2.22% 0.40% 1.52% 2.49% 16.3%  220 g/L  IPC residue of May 13,1994 7.82 4.511.74  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: 0  leach filtrate 24.1 14.4 25.4 0.99 14.8  g/L g/L g/L g/L g/L  0.379 M 0.244 M 0.434 M 0.013 M 0.265 M  66.8 g/L  0.681 M  repulp filtrate 1.07 0.48 0.80 0.036 0.54  g/L g/L g/L g/L g/L  13 g/L  leach residue 27.31 4.17 6.55  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 24.1  49.6% 78.3% 49.5% 74.0% 229.1  calc'd head grade 55.96% 7.57% 8.60% 0.60% 5.19%  calc'd head extraction 50.1% 77.8% 43.0% 74.9%  balance errors (w.r.t. head grades) absolute relative 0.65% 0.36% 0.08% 1.01% 0.19% 2.18% 0.08% 11.39% 3.64% 0.18%  Appendix C AUTOCLAVE TEST  test ID: series ID: date:  134  Experiment Worksheets WORKSHEET  #36 Kinetics, High Cu : S Wednesday, February 1, 1995 284 °F  140 °C 44 min 300 min  temperature: heat-up time to T: leaching time at T:  3.6 bar 3.6 bar  52 psig 52 psig  13.7 1/s  820 rpm  initial pressure: final pressure: stirring rate:  5 h  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  POTENTIAL  leach residue:  312 g  residue ratio:  leach filtrate: wash solution: total filtrate:  950 mL 812 mL 1762 mL  wash acidity:  repulp volume:  559 mL  1.15 wt./wt.  12 g / L H 2 S 0 4  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  63 mV (SCE)  308 mV (SHE)  30 mV (SCE) 130 mV (SCE)  275 mV (SHE) 375 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 13, 1994 56.32% 7.49% 8.79% 0.68% 5.01% 7.20% 12.50%  leach residue 74.0% 0.84% 0.60% 0.40% 0.34% 1.87% 16.3%  220 g/L  IPC residue of May 13,1994 7.82 4.51 1.74  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 0.3 25.3 30.9 1.24 17.8  g/L g/L g/L g/L g/L  0.004 M 0.430 M 0.527 M 0.017 M 0.319 M  69.3 g/L  0.707 M  repulp filtrate 0.26 0.47 0.56 0.028 0.37  g/L g/L g/L g/L g/L  12.5 g/L  leach residue 39.57 4.54 8.72  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 86.8  90.3% 94.1% 49.0% 94.1% 217.8  calc'd head grade 55.45% 7.41% 8.68% 0.67% 5.00%  calc'd head extraction 90.2% 94.0% 48.6% 94.1%  balance errors (w.r.t. head grades) absolute relative 0.87% 1.54% 0.08% 1.09% 0.11% 1.22% 0.01% 0.88% 0.01% 0.16%  Appendix C  135  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #35 Kinetics, High Cu : S Tuesday, January 31, 1995 284 °F  temperature: heat-up time to T: leaching time at T:  140 °C 44 min 0 min  initial pressure: final pressure:  3.4 bar 3.4 bar  50 psig 50 psig  13.7 1/s  820 rpm  stirring rate:  0 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  PdTENTIAL  360 g 700 mL 500 mL  leach residue:  321 g  residue ratio:  leach filtrate: wash solution: total filtrate:  871 mL 701 mL 1572 mL  wash acidity:  repulp volume:  480 mL  1.12 wt./wt.  14 g/L H2S04  MEASUREMENTS  before:  slurry:  65 mV (SCE)  310 mV (SHE)  after:  slurry: supernatant:  67 mV (SCE) 66 mV (SCE)  312 mV (SHE) 311 mV(SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 13,1994 56.32% 7.49% 8.79% 0.68% 5.01% 7.20% 12.50%  leach residue 64.0% 6.20% 0.32% 0.52% 1.84% 3.44% 16.0%  220 g/L  element ratio (wt./wt.) Cu: O C u : Stot Stot:0  IPC residue of May 13,1994 7.82 4:51 1.74  leach filtrate 32.3 10.9 25.1 0.85 14.4  g/L g/L g/L g/L g/L  0.508 M 0.184 M 0.428 M 0.011 M 0.258 M  82.7 g/L  0.843 M  repulp filtrate 2.85 0.39 0.99 0.040 0.59  g/L g/L g/L g/L g/L  16 g/L  leach residue 18.60 4.00 4.65  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 9.5  26.2% 96.8% 31.8% 67.3% 200.4  calc'd head grade 56.12% 8.16% 6.36% 0.67% 5.13%  calc'd head extraction 32.2% 95.5% 30.8% 68.0%  balance errors (w.r.t. head grades) absolute relative 0.20% 0.36% 0.67% 8.90% 2.43% 27.60% 0.01% 1.49% 0.12% 2.30%  Appendix C  Experiment Worksheets  136  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #34 Kinetics, High Cu : S Thursday, January 26, 1995  temperature: heat-up time to T: leaching time at T:  140 °C 45 min 30 min  initial pressure: final pressure:  3.8 bar 3.8 bar  stirring rate:  13.7 1/s  284 °F 0.5 h 55 psig 55 psig 820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach residue:  313 g  residue ratio:  leach filtrate: wash solution: total filtrate:  950 mL 548 mL 1498 mL  wash acidity:  repulp volume:  504 mL  1.15 wt./wt.  14 g / L H 2 S 0 4  P6TENTIAL M E A S U R E M E N T S before:  slurry:  61 mV (SCE)  306 mV (SHE)  after:  slurry: supernatant:  56 mV (SCE) 63 mV (SCE)  301 mV (SHE) 308 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 13,1994 56.32% 7.49% 8.79% 0.68% 5.01% 7.20% 12.50%  leach residue 68.0% 3.80% 1.78% 0.56% 1.24% 3.17% 16.4%  220 g/L  IPC residue of May 13,1994 7.82 4.51 1.74  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate 18.1 16.4 27.8 0.93 14.7  g/L g/L g/L g/L g/L  0.285 M 0.277 M 0.473 M 0.012 M 0.263 M  repulp filtrate 1.73 g/L 1.10 g/L 1.80 g/L 0.083 g/L 0.98 g/L  66.1 g/L  0.674 M  16 g/L  leach residue 21.45 4.15 5.17  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 39.7  55.9% 82.4% 28.4% 78.5% 232.1  calc'd head grade 55.15% 7.62% 8.87% 0.73% 4.96%  calc'd head extraction 56.6% 82.6% 33.4% 78.3%  balance errors (w.r.t. head grades) absolute relative 1.17% 2.08% 0.13% 1.72% 0.08% 0.95% 0.05% 7.58% 0.05% 0.99%  Appendix C  Experiment Worksheets  137  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #33 Kinetics, High Cu : S Wednesday, January 25, 1995 284 °F  temperature: heat-up time to T: leaching time at T:  140 °C 43 min 60 min  initial pressure: final pressure:  3.8 bar 3.8 bar  55 psig 55 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach residue:  314 g  residue ratio:  leach filtrate: wash solution: total filtrate:  949 mL 540 mL 1489 mL  wash acidity:  repulp volume:  626 mL  1.15 wt./wt.  14 g/L H2SQ4  P6TENTIAL M E A S U R E M E N T S before:  slurry:  64 mV (SCE)  309 mV (SHE)  after:  slurry: supernatant:  58 mV (SCE) 51 mV (SCE)  303 mV (SHE) 296 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 13, 1994 56.32% 7.49% 8.79% 0.68% 5.01% 7.20% 12.50%  leach residue 70.0% 2.60% 1.36% 0.48% 1.06% 2.47% 16.3%  220 g/L  IPC residue of May 13, 1994 7.82 4.51 1.74  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 12.8 19.4 28.3 1.23 15.7  g/L g/L g/L g/L g/L  0.201 M 0.330 M 0.483 M 0.016 M 0.281 M  66.6 g/L  0.679 M  repulp filtrate 1.06 0.90 1.35 0.075 0.75  g/L g/L g/L g/L g/L  15.5 g/L  leach residue 28.34 4.29 6.60  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 53.8  69.7% 86.5% 38.4% 81.5% 231.3  calc'd head grade 55.67% 7.39% 8.66% 0.74% 5.07%  calc'd head extraction 69.3% 86.3% 43.6% 81.8%  balance errors (w.r.t. head grades) absolute relative 0.65% 1.15% 0.10% 1.37% 0.13% 1.49% 0.06% 9.11% 0.06% 1.14%  Appendix C  Experiment Worksheets  138  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #32 Kinetics, High Cu : S Saturday, September 3,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 51 min 0 min  initial pressure: final pressure:  5.5 bar 5.5 bar  80 psig 80 psig  13.7 1/s  820 rpm  stirring rate:  0 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  leach residue:  360 g 700 mL 500 mL  leach filtrate: wash water: total filtrate: repulp volume:  308 g  residue ratio:  975 mL 654 mL 1629 mL  wash acidity:  1.17 wt./wt.  0 g/L H2S04  545 mL  P6TENTIAL MEASUREMENTS before:  slurry:  62 mV (SCE)  307 mV (SHE)  after:  slurry: supernatant:  64 mV (SCE) 70 mV (SCE)  309 mV (SHE) 315 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 12,1994 57.89% 7.44% 8.63% 0.70% 4.79% 6.33% 13.37%  leach residue 69.9% 4.30% 2.14% 0.52% 1.47% 3.4% 15.9%  220 g/L  IPC residue of May 12,1994 9.15 4.33 2.11  element ratio (wt./wt.) Cu : 0 C u : Stot Stot: O  leach filtrate 21.7 15.6 24.4 0.98 14.9  g/L g/L g/L g/L g/L  0.341 M 0.264 M 0.416 M 0.013 M 0.267 M  69.1 g/L  0.704 M  repulp filtrate 0.537 0.313 0.635 0.025 0.300  g/L g/L g/L g/L g/L  1 g/L  leach residue 20.56 4.40 4.68  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 28.8  50.6% 78.8% 36.8% 73.7% 240.7  calc'd head grade 56.93% 7.90% 8.44% 0.71% 5.29%  calc'd head extraction 53.4% 78.3% 37.6% 76.2%  balance errors (w.r.t. head grades) absolute relative 0.96% 1.66% 0.46% 6.17% 0.19% 2.17% 0.01% 1.22% 0.50% 10.45%  Appendix C Experiment Worksheets AUTOCLAVE TEST  test ID: series ID: date:  139  WORKSHEET  #31 Kinetics, High Cu : S Friday, September 2,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 51 min 15 min  initial pressure: final pressure:  5.5 bar 5.5 bar  80 psig 80 psig  13.7 1/s  820 rpm  stirring rate:  0.25 h  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  308 g  residue ratio:  999 mL 630 mL 1629 mL  wash acidity:  1.17 wt./wt.  0 g/L H2S04  517 mL  MEASUREMENTS  before:  slurry:  62 mV (SCE)  307 mV (SHE)  after:  slurry: supernatant:  55 mV (SCE) 70 mV (SCE)  300 mV (SHE) 315 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 12, 1994 57.89% 7.44% 8.63% 0.70% 4.79% 6.33% 13.37%  leach residue 74.3% 2.45% 1.19% 0.46% 0.82% 3.4% 16.0%  69.0 g/L  220 g/L  IPC residue of May 12,1994 9.15 4.33 2.11  element ratio (wtVwt.) Cu:0 C u : Stot Stot:0  leach filtrate 10.6 g/L . 20.4 g/L 25.8 g/L 0.95 g/L 15.6 g/L  0.167 0.346 0.440 0.013 0.279  M M M M M  0.704 M  repulp filtrate 0.709 g/L 0.465 g/L 0.411 g/L 0.035 g/L 0.437 g/L  1 g/L  leach residue 21.85 4.64 4.71  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 58.1  71.8% 88.2% 43.3% 85.3% 236.3  calc'd head grade 57.76% 7.75% 8.18% 0.66% 5.03%  calc'd head extraction 73.0% 87.6% 39.8% 86.0%  balance errors (w.r.t. head grades) absolute relative 0.13% 0.23% 0.31% 4.18% 0.45% 5.20% 0.04% 5.72% 0.24% 5.09%  Appendix C  Experiment Worksheets  140  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #30 Kinetics, High Cu : S Thursday, September 1,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 50 min 30 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  stirring rate:  0.5 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  leach residue:  360 g 700 mL 500 mL  leach filtrate: wash water: total filtrate: repulp volume:  309 g  residue ratio:  942 mL 630 mL 1572 mL  wash acidity:  1.17 wt./wt.  0 g/L H2S04  452 mL  P6TENTIAL M E A S U R E M E N T S before:  slurry:  67 mV (SCE)  312 mV (SHE)  after:  slurry: supernatant:  44 mV (SCE) 56 mV (SCE)  289 mV (SHE) 301 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 12,1994 57.89% 7.44% 8.63% 0.70% 4.79% 6.33% 13.37%  leach residue 74.0% 1.58% 0.89% 0.43% 0.67% 2.7% 16.8%  220 g/L  IPC residue of May 12,1994 9.15 4.33 2.11  element ratio (wt./wt.) Cu:0 C u : Stot Stot: O  leach filtrate 5.97 20.4 30.9 1.34 18.8  g/L g/L g/L g/L g/L  0.094 M 0.345 M 0.526 M 0.018 M 0.337 M  72.7 g/L  0.741 M  repulp filtrate 0.896 0.687 0.330 0.049 0.551  g/L g/L g/L g/L g/L  2 g/L  leach residue 27.41 4.40 6.22  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 71.9  81.8% 91.1% 47.8% 88.0% 237.5  calc'd head grade 56.33% 6.68% 8.84% 0.72% 5.50%  calc'd head extraction 79.7% 91.3% 49.0% 89.5%  balance errors (w.r.t. head grades) absolute relative 1.56% 2.70% 0.76% 10.18% 0.21% 2.47% 0.02% 2.40% 0.71% 14.79%  Appendix C  141  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #29 Kinetics, High Cu : S Wednesday, August 31,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 48 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  stirring rate:  1h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  leach residue:  360 g 700 mL 500 mL  leach filtrate: wash water: total filtrate: repulp volume:  310 g  residue ratio:  955 mL 596 mL 1551 mL  wash acidity:  1.16 wtVwt.  0 g/L H2S04  518 mL  P6TENTIAL M E A S U R E M E N T S before:  slurry:  after:  slurry: supernatant:  69 mV (SCE)  314 mV (SHE)  20 mV (SCE) 200 mV (SCE)  265 mV (SHE) 445 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 12, 1994 57.89% 7.44% 8.63% 0.70% 4.79% 6.33% 13.37%  leach residue 77.1% 0.78% 0.60% 0.41% 0.38% 4.0% 16.4%  220 g/L  IPC residue of May 12,1994 9.15 4.33 2.11  element ratio (wt./wt.) Cu:0 Cu : Stot Stot: O  leach filtrate 1.18 22.4 28.6 1.31 19.2  g/L g/L g/L g/L g/L  0.019 M 0.381 M 0.487 M 0.017 M 0.345 M  70.4 g/L  0.718 M  repulp filtrate 0.857 0.667 0.027 0.046 0.534  g/L g/L g/L g/L g/L  1 g/L  leach residue 19.28 4.70 4.10  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 84.4  90.9% 94.0% 49.9% 93.3% 241.1  calc'd head grade 58.11% 6.63% 8.10% 0.70% 5.43%  calc'd head extraction 89.8% 93.6% 49.7% 94.0%  balance errors (w.r.t. head grades) absolute relative 0.22% 0.37% 0.81% 10.89% 0.53% 6.08% 0.00% 0.36% 0.64% 13.30%  Appendix C  Experiment Worksheets  AUTOCLAVE TEST  test ID: series ID: date:  142  WORKSHEET  #28 Kinetics, High Cu : S Tuesday, August 30, 1994 160 °C 51 min 240 min  temperature: heat-up time to T: leaching time at T:  4 h  5.9 bar 8.3 bar  85 psig 120 psig  13.7 1/s  820 rpm  initial pressure: final pressure: stirring rate:  320 °F  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  leach residue:  360 g 700 mL 513 mL  leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  300 g  residue ratio:  980 mL 689 mL 1669 mL  wash acidity:  1.20 wt./wt.  0 g/L H2S04  656 mL  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  69 mV (SCE)  314 mV (SHE)  -160 mV (SCE) 60 mV (SCE)  85mV(SHE) 305 mV (SHE)  ASSAYS  component Cu Co Ni As Fe 0 Stot H2S04  electrolyte 45 g/L  IPC residue of May 12,1994 57.89% 7.44% 8.63% 0.70% 4.79% 6.33% 13.37%  leach residue 76.2% 0.06% 0.90% 0.73% 0.11% 4.4% 16.6%  220 g/L  element ratio (wt./wt.) Cu:0 C u : Stot Stot:0  IPC residue of May 12,1994 9.15 4.33 2.11  leach filtrate 0.62 g/L 19.5 g/L 27.2 g/L Og/L 12.2 g/L  0.010 M 0.331 M 0.464 M 0M 0.218 M  repulp filtrate 0.924 g/L 0.665 g/L Og/L Og/L 0.619 g/L  60.3 g/L  0.615 M  1 g/L  leach residue 17.32 4.59 3.77  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 85.8  99.4% 91.3% 12.7% 98.2% 263.7  calc'd head grade 54.92% 5.36% 8.16% 0.61% 3.41%  calc'd head extraction 99.1% 90.9% 0.1% 97.4%  balance errors (w.r.t. head grades) absolute relative 2.97% 5.13% 2.08% 27.95% 0.47% 5.39% 0.09% 12.69% 1.38% 28.84%  Appendix C  Experiment Worksheets  143  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #27 Pulp Density Friday, August 26, 1994  temperature: heat-up time to T: leaching time at T:  160 °C 50 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  320 1 85 psig 85 psig  13.7 1/s 29.5% solids  stirring rate: pulp density:  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed 617.1 g 1200 mL 0 mL  IPC residue: electrolyte: deionized water:  leach residue:  485.6 g  leach filtrate: wash water: total filtrate: repulp volume:  PfitENTlAL  residue ratio:  871 mL 751 mL 1622 mL  wash acidity:  1.27 wt./wt.  0 g/L H2S04  687 mL  MEASUREMENTS  before:  slurry:  95 mV (SCE)  340 mV (SHE)  after:  slurry: supernatant:  85 mV (SCE) mV (SCE)  330 mV (SHE) mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 22,1994 59.17% 6.13% 7.76% 0.68% 4.07% 5.98% 14.86%  leach residue 74.3% 0.39% 0.37% 0.27% 0.28% 4.7% 19.4%  220 g/L IPC residue of May 22,1994 9.89 3.98 2.48  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 50.5 43.1 47.1 3.38 30.5  g/L g/L g/L g/L g/L  0.794 M 0.731 M 0.803 M 0.045 M 0.547 M  91.0 g/L  0.927 M  repulp filtrate 1.201 0.916 1.205 0.085 0.709  g/L g/L g/L g/L g/L  2 g/L  leach residue 15.81 3.83 4.13  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 16.3  95.0% 96.2% 69.3% 94.5% 299.4  calc'd head grade 56.84% 6.39% 6.94% 0.69% 4.53%  calc'd head extraction 95.2% 95.8% 69.6% 95.1%  balance errors (w.r.t. head grades) absolute relative 2.33% 3.93% 0.26% 4.26% 0.82% 10.52% 0.01% 0.86% 0.46% 11.39%  Appendix C  144  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #26 Pulp Density Thursday, August 25, 1994  temperature: heat-up time to T: leaching time at T:  160 °C 53 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  320 °F 1 h 85 psig 85 psig  13.7 1/s 15.0% solids  stirring rate: pulp density:  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  leach residue:  229.9 g 447 mL 753 mL  leach filtrate: wash water: total filtrate: repulp volume:  198.2 g  residue ratio:  1049 mL 629 mL 1678 mL  wash acidity:  1.16 wt./wt.  0 g/L H2S04  500 mL  P6TENTIAL M E A S U R E M E N T S before:  slurry:  63 mV (SCE)  308 mV (SHE)  after:  slurry: supernatant:  34 mV (SCE) 70 mV (SCE)  279 mV (SHE) 315 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 22, 1994 59.17% 6.13% 7.76% 0.68% 4.07% 5.98% 14.86%  leach residue 77.2% 0.37% 0.45% 0.36% 0.23% 5.1% 17.2%  220 g/L IPC residue of May 22, 1994 9.89 3.98 2.48  element ratio (wt./wt.) Cu:0 Cu : Stot Stot:O  leach filtrate 0.25 14.1 17.0 0.61 9.2  g/L g/L g/L g/L g/L  0.004 M 0.239 M 0.289 M 0.008 M 0.164 M  46.4 g/L  0.473 M  repulp filtrate 0.124 0.019 0.043 0.002 0.122  g/L g/L g/L g/L g/L  0.1 g/L  leach residue 15.14 4.49 3.37  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 86.4  94.8% 95.0% 54.1% 95.1% 215.9  calc'd head grade 57.92% 6.74% 8.14% 0.59% 4.38%  calc'd head extraction 95.3% 95.2% 47.1% 95.5%  balance errors (w.r.t. head grades) absolute relative 1.25% 2.12% 0.61% 10.02% 0.38% 4.84% 0.09% 13.25% 0.31% 7.72%  Appendix C AUTOCLAVE TEST  test ID: series ID: date:  145  Experiment Worksheets WORKSHEET  #25 Pulp Density Wednesday, August 24,1994  temperature: heat-up time to T: leaching time at T:  160 °C 53 min 60 min  initial pressure: final pressure:  6.1 bar 6.1 bar  stirring rate: pulp density:  320 °F 1h 88 psig 88 psig  13.7 1/s 10.0% solids  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  140.3 g 273 mL 927 mL  IPC residue: electrolyte: deionized water:  leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  122.1 g  residue ratio:  1096 mL 476 mL 1572 mL  wash acidity:  1.15 wt./wt.  0 g/L H2S04  496 mL  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  51 mV (SCE)  296 mV (SHE)  70 mV (SCE) 125 mV (SCE)  315 mV (SHE) 370 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 22, 1994 59.17% 6.13% 7.76% 0.68% 4.07% 5.98% 14.86%  leach residue 75.9% 0.38% 0.46% 0.32% 0.23% 6.1% 17.2%  220 g/L IPC residue of May 22,1994 9.89 3.98 2.48  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 1.24 9.2 9.2 0.52 6.2  g/L g/L g/L g/L g/L  0.019 M 0.156 M 0.156 M 0.007 M 0.112 M  25.9 g/L  0.264 M  repulp filtrate 0.197 0.112 0.105 0.005 0.077  g/L g/L g/L g/L g/L  0.1 g/L  leach residue 12.44 4.41 2.82  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 77.9  94.6% 94.8% 59.4% 95.2% 226.0  calc'd head grade 58.26% 7.51% 7.57% 0.68% 5.07%  calc'd head extraction 95.6% 94.7% 59.5% 96.1%  balance errors (w.r.t. head grades) absolute relative 0.91% 1.53% 1.38% 22.49% 0.19% 2.50% 0.00% 0.15% 1.00% 24.46%  Appendix C  Experiment Worksheets  146  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #24 Pulp Density Tuesday, August 23,1994  temperature: heat-up time to T: leaching time at T:  160 °C 55 min 60 min  initial pressure: final pressure:  6.2 bar 6.2 bar  320 °F 1 h 90 psig 90 psig  13.7 1/s 5.0% solids  stirring rate: pulp density:  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed 64.7 g 126 mL 1074 mL  IPC residue: electrolyte: deionized water:  leach residue:  59.2 g  residue ratio:  leach filtrate: wash water: total filtrate:  1141 mL 314 mL 1455 mL  wash acidity:  repulp volume:  1.09 wt./wt.  0 g/L H2S04  486 mL  P6TENTIAL MEASUREMENTS before:  slurry:  after:  slurry: supernatant:  56 mV (SCE)  301 mV (SHE)  70 mV (SCE) 104 mV (SCE)  315 mV (SHE) 349 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 22,1994 59.17% 6.13% 7.76% 0.68% 4.07% 5.98% 14.86%  leach residue 76.1% 0.36% 0.45% 0.32% 0.25% 4.0% 18.2%  220 g/L IPC residue of May 22,1994 9.89 3.98 2.48  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate 0.78 4.0 4.2 0.23 2.5  g/L g/L g/L g/L g/L  0.012 M 0.068 M 0.072 M 0.003 M 0.044 M  10.2 g/L  0.104 M  repulp filtrate 0.242 0.131 0.152 0.007 0.184  g/L g/L g/L g/L g/L  0.1 g/L  leach residue 19.03 4.18 4.55  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 73.9  94.6% 94.7% 57.3% 94.4% 247.8  calc'd head grade 62.24% 7.40% 7.90% 0.70% 4.55%  calc'd head extraction 95.5% 94.8% 58.6% 95.0%  balance errors (w.r.t. head grades) absolute relative 3.07% 5.19% 1.27% 20.76% 0.14% 1.83% 0.02% 2.96% 0.48% 11.74%  Appendix C AUTOCLAVE TEST  test ID: series ID: date:  147  Experiment Worksheets WORKSHEET  #23 Pulp Density Friday, August 19, 1994  temperature: heat-up time to T: leaching time at T:  160 °C 52 min 60 min  initial pressure: final pressure:  6.1 bar 6.1 bar  320 °F 1 h 88 psig 88 psig  13.7 1/s 20.9% solids  stirring rate: pulp density:  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360.0 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  308.9 g  leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  residue ratio:  984 mL 756 mL 1740 mL  wash acidity:  1.17 wt./wt.  0 g/L H2S04  470 mL  MEASUREMENTS  before:  slurry:  after:  slurry: supernatant:  74mV(SCE)  319 mV (SHE)  68 mV (SCE) 140 mV (SCE)  313 mV (SHE) 385 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 22, 1994 59.17% 6.13% 7.76% 0.68% 4.07% 5.98% 14.86%  leach residue 77.0% 0.33% 0.42% 0.38% 0.24% 5.4% 17.6%  220 g/L IPC residue of May 22,1994 9.89 3.98 2.48  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate 3.00 22.2 19.2 1.19 15.6  g/L g/L g/L g/L g/L  0.047 M 0.376 M 0.327 M 0.016 M 0.279 M  79.6 g/L  0.812 M  repulp filtrate 0.240 0.127 0.149 0.006 0.187  g/L g/L g/L g/L g/L  0.1 g/L  leach residue 14.26 4.38 3.26  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 79.3  95.3% 95.4% 52.3% 94.9% 210.1  calc'd head grade 58.14% 6.34% 5.60% 0.65% 4.47%  calc'd head extraction 95.5% 93.6% 50.0% 95.4%  balance errors (w.r.t. head grades) absolute relative 1.03% 1.74% 0.21% 3.50% 2.16% 27.86% 0.03% 4.56% 0.40% 9.73%  Appendix C  148  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #22 Agitation Thursday, August 18, 1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 51 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  10.4 1/s  625 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach filtrate: wash water: total filtrate: repulp volume:  297 g  residue ratio:  979 mL 527 mL 1506 mL  wash acidity:  1.21 wt./wt.  0 g/L H2S04  661 mL  P6TENTIAL M E A S U R E M E N T S before:  slurry:  after:  supernatant ?:  70 mV (SCE)  315 mV (SHE)  228 mV (SCE)  473 mV (SHE)  ASSAYS  component Cu Co Ni As Fe 0 Stot H2S04  electrolyte 45 g/L  IPC residue of May 18, 1994 54.80% 7.47% 8.98% 0.77% 4.84% 6.60% 14.51%  leach residue 74.8% 0.89% 0.63% 0.74% 0.36% 4.3% 17.5%  220 g/L IPC residue of May 18, 1994 8.30 3.78 2.20  element ratio (wt./wt.) Cu:0 Cu : Stot Stot:0  leach filtrate 0.48 26.6 30.1 0.24 18.8  g/L g/L g/L g/L g/L  0.008 M 0.451 M 0.512 M 0.003 M 0.337 M  repulp filtrate 0.706 g/L 0.520 g/L 0.119 g/L 0.056 g/L 0.458 g/L  66.8 g/L  0.681 M  1 g/L  leach residue 17.40 4.27 4.07  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 86.2  90.1% 94.2% 21.0% 93.8% 246.1  calc'd head grade 53.09% 7.97% 8.70% 0.67% 5.41%  calc'd head extraction 90.7% 94.0% 9.6% 94.5%  balance errors (w.r.t. head grades) relative absolute 3.12% 1.71% 0.50% 6.63% 3.13% 0.28% 12.64% 0.10% 11.78% 0.57%  Appendix C  Experiment Worksheets  149  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #21 Agitation Wednesday, August 17,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 50 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  16.7 1/s  1000 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  360 g 700 mL 500 mL  leach residue:  . 307 g  residue ratio:  leach filtrate: wash water: total filtrate:  974 mL 666 mL 1640 mL  wash acidity:  repulp volume: P6TENTIAL  1.17 wt./wt.  0 g/L H2S04  702 mL  MEASUREMENT  before:  slurry:  after:  supernatant?:  74 mV (SCE)  319 mV (SHE)  157 mV (SCE)  402 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O  electrolyte 45 g/L  Stot H2S04  IPC residue of May 18,1994 54.80% 7.47% 8.98% 0.77% 4.84% 6.60% 14.51%  leach residue 73.8% 0.90% 0.78% 0.81% 0.40% 5.4% 17.8%  220 g/L  element ratio (wt./wt.) Cu : 0 C u : Stot Stot:O  IPC residue of May 18,1994 8.30 3.78 2.20  leach residue 13.67 4.15 3.30  consumption kg/t IPC residue 86.3  calc'd head grade 54.31% 8.09% 8.82% 0.70% 5.72%  leach filtrate 0.45 27.1 30.1 0.01 19.9  g/L g/L g/L g/L g/L  0.007 M 0.459 M 0.514 M 0.000 M 0.356 M  64.1 g/L  0.653 M  repulp filtrate 0.625 0.391 0.003 0.004 0.368  g/L g/L g/L g/L g/L  0.1 g/L  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction 89.7% 92.6% 9.8% 93.0%  254.5  calc'd head extraction 90.5% 92.5% 0.4% 94.1%  balance errors (w.r.t. head grades) absolute relative 0.49% 0.90% 0.62% 8.26% 0.16% 1.80% 0.07% 9.51% 0.88% 18.10%  Appendix C  150  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #20 Agitation Tuesday, August 16,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 50 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  14.6 1/s  875 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  298 g  residue ratio:  976 mL 872 mL 1848 mL  wash acidity:  1.21 wt./wt.  0 g/L H2S04  452 mL  MEASUREMENTS  before:  slurry:  after:  supernatant?:  76 mV (SCE)  321 mV (SHE)  197 mV (SCE)  442 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 18,1994 54.80% 7.47% 8.98% 0.77% 4.84% 6.60% 14.51%  leach residue 74.2% 0.91% 0.64% 0.44% 0.43% 5.4% 17.3%  220 g/L IPC residue of May 18,1994 8.30 3.78 2.20  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 0.68 24.3 30.4 1.50 19.6  g/L g/L g/L g/L g/L  0.011 M 0.413 M 0.517 M 0.020 M 0.350 M  70.5 g/L  0.719 M  repulp filtrate 0.371 0.238 0.142 0.011 0.206  g/L g/L g/L g/L g/L  1 g/L  leach residue 13.74 4.29 3.20  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 85.7  89.9% 94.1% 53.2% 92.7% 236.6  calc'd head grade 52.85% 7.35% . 8.76% 0.77% 5.66%  calc'd head extraction 89.7% 94.0% 53.1% 93.8%  balance errors (w.r.t. head grades) relative absolute 1.95% 3.55% 0.12% 1.54% 0.22% 2.48% 0.00% 0.39% 0.82% 16.88%  Appendix C  151  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #19 Agitation Friday, August 12,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 54 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  12.5 1/s  750 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  leach residue:  360 g 700 mL 500 mL  leach filtrate: wash water: total filtrate: repulp volume:  295 g  residue ratio:  827 mL 1063 mL 1890 mL  wash acidity:  1.22 wt./wt.  0 g/L H2S04  670 mL  POTENTIAL MEASUREMENT S before:  slurry:  after:  supernatant?:  94 mV (SCE)  339 mV (SHE)  177 mV (SCE)  422 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 18,1994 54.80% 7.47% 8.98% 0.77% 4.84% 6.60% 14.51%  leach residue 75.4% 0.84% 0.61% 0.69% 0.41% 3.2% 18.0%  220 g/L IPC residue of May 18,1994 8.30 3.78 2.20  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 0.17 30.4 33.3 0.30 22.1  g/L g/L g/L g/L g/L  0.003 M 0.516 M 0.567 M 0.004 M 0.396 M  77.7 g/L  0.792 M  repulp filtrate 0.125 0.290 0.470 0.030 0.133  g/L g/L g/L g/L g/L  Og/L  leach residue 23.56 4.19 5.63  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 87.1  90.8% 94.5% 26.3% 93.0% 249.3  calc'd head grade 53.08% 7.67% 8.14% 0.64% 5.42%  calc'd head extraction 91.0% 93.9% 10.8% 93.8%  balance errors (w.r.t. head grades) absolute relative 1.72% 3.15% 0.20% 2.72% 0.84% 9.30% 0.13% 17.34% 0.58% 12.04%  Appendix C  152  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #18 Acid Thursday, August 11,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 52 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  303 g  residue ratio:  973 mL 796 mL 1769 mL  wash acidity:  1.19 wt./wt.  0 g/L H2S04  548 mL  MEASUREMENTS  before:  slurry:  69 mV (SCE)  314 mV (SHE)  after:  slurry:  132 mV (SCE)  377 mV (SHE)  ASSAYS  component Cu Co Ni As Fe 0 Stot H2S04  electrolyte 45 g/L  IPC residue of May 9,1994 54.94% 6.76% 9.54% 0.71% 4.11% 7.94% 15.38%  leach residue 74.0% 0.39% 1.06% 0.83% 0.21% 2.8% 18.2%  220 g/L IPC residue of May 9,1994 6.92 3.57 1.94  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: 0  leach filtrate 0.1 25.2 31.8 0.01 17.5  g/L g/L g/L g/L g/L  0.002 M 0.428 M 0.542 M 0.000 M 0.313 M  repulp filtrate 0.207 g/L 0.028 g/L 0g/L 0.001 g/L 0.119 g/L  69.1 g/L  0.705 M  0.1 g/L  leach residue 26.43 4.07 6.50  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 87.2  95.2% 90.6% 1.8% 95.7% 240.9  calc'd head grade 53.56% 7.14% 9.49% 0.70% 4.90%  calc'd head extraction 95.4% 90.6% 0.4% 96.4%  balance errors (w.r.t. head grades) relative absolute 2.50% 1.38% 5.58% 0.38% 0.51% 0.05% 1.48% 0.01% 19.14% 0.79%  Appendix C  Experiment Worksheets  AUTOCLAVE TEST  test ID: series ID: date:  153  WORKSHEET  #17 Acid Wednesday, August 10,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 51 min 60 min  initial pressure: final pressure:  5.9 bar 5.9 bar  85 psig 85 psig  13.7 1/s  820 rpm  stirring rate:  1h  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  306 g  residue ratio:  981 mL 869 mL 1850 mL  wash acidity:  1.18 wt./wt.  0 g/L H2S04  445 mL  MEASUREMENTS  before:  slurry:  69 mV (SCE)  314 mV (SHE)  after:  slurry:  50 mV (SCE)  295 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 9, 1994 54.94% 6.76% 9.54% 0.71% 4.11% 7.94% 15.38%  leach residue 74.5% 0.42% 1.13% 0.81% 0.25% 3.0% 18.0%  150 g/L IPC residue of May 9, 1994 6.92 3.57 1.94  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate 0.2 25.8 29.6 0.01 16.2  g/L g/L g/L g/L g/L  0.003 M 0.437 M 0.504 M 0.000 M 0.291 M  repulp filtrate 0.363 g/L 0.121 g/L Og/L 0.002 g/L 0.164 g/L  22.7 g/L  0.231 M  0.1 g/L  leach residue 24.83 4.14 6.00  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 87.0  94.8% 89.9% 3.1% 94.7% 229.9  calc'd head grade 54.62% 7.37% 9.03% 0.69% 4.64%  calc'd head extraction 95.2% 89.4% 0.3% 95.3%  balance errors (w.r.t. head grades) relative absolute 0.58% 0.32% 9.05% 0.61% 5.36% 0.51% 2.82% 0.02% 12.90% 0.53%  Appendix C  154  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #16 Acid Tuesday, August 9, 1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 51 min 60 min  initial pressure: final pressure:  6.2 bar 6.2 bar  90 psig 90 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach filtrate: wash water: total filtrate: repulp volume:  319 g  residue ratio:  956 mL 604 mL 1560 mL  wash acidity:  1.13 wt./wt.  0 g/L H2S04  530 mL  P6TENTIAL MEASUREMENTS before:  slurry:  64 mV (SCE)  309 mV (SHE)  after:  slurry:  78 mV (SCE)  323 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 9,1994 54.94% 6.76% 9.54% 0.71% 4.11% 7.94% 15.38%  leach residue 74.5% 0.51% 2.01% 0.85% 0.50% 2.7% 18.1%  100 g/L IPC residue of May 9,1994 6.92 3.57 1.94  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot : O  leach filtrate 0.1 26.6 27.6 0.01 15.3  g/L g/L g/L g/L g/L  0.001 M 0.452 M 0.470 M 0.000 M 0.274 M  0.2 g/L  0.002 M  repulp filtrate 0.127 0.942 0 0.001 0.590  g/L g/L g/L g/L g/L  0.1 g/L  leach residue 27.59 4.12 6.70  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 87.3  93.3% 81.3% -6.0% 89.3% 193.9  calc'd head grade 57.28% 7.53% 9.10% 0.75% 4.51%  calc'd head extraction 93.9% 80.4% 0.3% 90.3%  balance errors (w.r.t. head grades) relative absolute 4.27% 2.34% 11.35% 0.77% 4.57% 0.44% 0.04% 6.28% 9.70% 0.40%  155  Appendix C Experiment Worksheets AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #15 Acid Friday, August 5,1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 53 min 60 min  initial pressure: final pressure:  6.2 bar 6.2 bar  90 psig 90 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360 g 700 mL 500 mL  IPC residue: electrolyte: deionized water:  leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  344 g  residue ratio:  995 mL 757 mL 1752 mL  wash acidity:  1.05 wt./wt.  0 g/L H2S04  557 mL  MEASUREMENTS  before:  slurry:  60 mV (SCE)  305 mV (SHE)  after:  slurry:  -317 mV (SCE)  -72 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 9,1994 54.94% 6.76% 9.54% 0.71% 4.11% 7.94% 15.38%  leach residue 67.0% 1.13% 5.81% 0.76% 1.51% 5.9% 17.0%  0.7 g/L  50 g/L IPC residue of May 9, 1994 6.92 3.57 1.94  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 0.1 g/L 22.1 g/L 15.6 g/L Og/L 9.6 g/L  0.001 M 0.374 M 0.265 M 0 M 0.171 M  0.008 M  repulp filtrate 0.111 g/L 1.270 g/L 0.001 g/L 0.002 g/L 0.580 g/L  1 g/L  leach residue 11.36 3.94 2.88  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 87.3  84.0% 41.8% -2.8% 64.9% 95.2  calc'd head grade 55.29% 7.18% 9.85% 0.73% 4.08%  calc'd head extraction 85.0% 43.7% 0.2% 64.7%  balance errors (w.r.t. head grades) absolute relative 0.64% 0.35% 0.42% 6.21% 0.31% 3.29% 0.02% 3.00% 0.03% 0.66%  156  Appendix C Experiment Worksheets AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #14 Acid Thursday, August 4,1994  temperature: heat-up time to T: leaching time at T:  160 °C 52 min 60 min  320 °F *  initial pressure: final pressure:  6.2 bar 6.2 bar  90 psig 90 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: deionized water:  leach residue:  360 g 700 mL 500 mL  leach filtrate: wash water: total filtrate: repulp volume:  residue ratio:  376 g 962 mL 678 mL 1640 mL  wash acidity:  0.96 wt./wt.  0 g/L H2S04  592 mL  P6TENTIAL M E A S U R E M E N T S before:  slurry:  61 mV (SCE)  306 mV (SHE)  after:  slurry:  -473 mV (SCE)  -228 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 9, 1994 54.94% 6.76% 9.54% 0.71% 4.11% 7.94% 15.38%  leach residue 60.7% 1.90% 8.43% 0.69% 2.89% 8.1% 15.5%  0.2 g/L  0 g/L  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate Og/L 20.1 g/L 3.1 g/L Og/L 4.1 g/L  IPC residue of May 9, 1994 6.92 3.57 1.94  0 M 0.342 M 0.052 M 0 M 0.073 M  0.002 M  repulp filtrate 0.001 g/L 0.175 g/L 0.104 g/L 0.002 g/L 0.008 g/L  0.1 g/L  leach residue 7.49 3.92 1.91  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 87.5  70.6% 7.7% -1.4% 26.6% -0.6  calc'd head grade 54.65% 7.37% 9.62% 0.72% 4.11%  calc'd head extraction 73.1% 8.5% 0.2% 26.6%  balance errors (w.r.t. head grades) relative absolute 0.53% 0.29% 8.96% 0.61% 0.08% 0.88% 1.53% 0.01% 0.00% 0.08%  Appendix C  Experiment Worksheets  157  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #12 Copper Wednesday, May 25,1994 320 °F  160 °C 54 min 60 min  temperature: heat-up time to T: leaching time at T:  1h  initial pressure: final pressure:  5.5 bar 16.5 bar  80 psig 240 psig  stirring rate:  13.7 1/s  820 rpm  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: distilled water:  leach residue: leach filtrate: wash water: total filtrate: repulp volume:  290 g  residue ratio:  1007 mL 873 mL 1880 mL  wash acidity:  1.24 wt./wt.  0 g/L H2SQ4  760 mL  P6TENTIAL M E A S U R E M E N T S before:  supernatant ?:  88 mV (SCE)  after:  supernatant ?:  197 mV (SCE)  333 mV (SHE) 48 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 15 g/L  IPC residue of May 20,1994 55.35% 7.09% 9.13% 0.76% 5.04% 6.23% 14.72%  leach residue 72.78% 1.31% 2.16% 0.95% 0.50% 2.60% 17.75%  220 g/L IPC residue of May 20,1994 8.88 3.76 2.36  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot : O  leach filtrate Og/L 24.2 g/L 29.5 g/L Og/L 18.4 g/L  0M 0.410 M 0.503 M 0M 0.329 M  52.3 g/L  0.533 M  repulp filtrate 0 0.017 0.153 0 0  g/L g/L g/L g/L g/L  0 g/L  leach residue 27.99 4.10 6.83  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2SQ4  assay head extraction  consumption kg/t IPC residue 29.2  85.1% 80.9% -0.7% 92.0% 281.6  calc'd head grade 55.71% 7.81% 9.99% 0.77% 5.54%  calc'd head extraction 86.5% 82.6% 0.0% 92.7%  balance errors (w.r.t. head grades) relative absolute 0.36% 0.65% 10.16% 0.72% 9.47% 0.86% 0.69% 0.01% 9.89% 0.50%  Appendix C  158  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #11 Copper Monday, May 23,1994  temperature: heat-up time to T: leaching time at T:  160 °C 114 min 60 min  320 °F  initial pressure: final pressure:  5.5 bar 7.2 bar  80 psig 105 psig  13.7 1/s  820 rpm  stirring rate:  heating and agitation problems 1 h  REAGENT BALANCE Autoclave products  Autoclave feed  leach residue:  360 g 700 mL 500 mL  IPC residue: electrolyte: distilled water:  leach filtrate: wash water: total filtrate: repulp volume:  294 g  residue ratio:  1020 mL 1010 mL 2030 mL  wash acidity:  1.22 wt./wt.  0 g/L H2S04  516 mL  P6TENTIAL M E A S U R E M E N T S before:  supernatant?:  101 mV(SCE)  346 mV (SHE)  after:  supernatant ?:  189 mV (SCE)  56 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 30 g/L  IPC residue of May 20, 1994 55.35% 7.09% 9.13% 0.76% 5.04% 6.23% 14.72%  leach residue 74.15% 1.03% 1.79% 0.92% 0.42% 2.58% 17.95%  220 g/L IPC residue of May 20, 1994 8.88 3.76 2.36  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate Og/L 24.5 g/L 29.7 g/L Og/L 18.3 g/L  0M 0.415 M 0.507 M 0M 0.328 M  repulp filtrate 0 g/L 0.056 g/L 0.226 g/L 0g/L Og/L  61.7 g/L  0.629 M  0 g/L  leach residue 28.74 4.13 6.96  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 58.3  88.1% 84.0% 1.1% 93.2% 253.0  calc'd head grade 54.72% 7.77% 9.89% 0.75% 5.53%  calc'd head extraction 89.2% 85.2% 0.0% 93.8%  balance errors (w.r.t. head grades) relative absolute 0.63% 1.13% 9.60% 0.68% 8.27% 0.76% 1.14% 0.01% 0.49% 9.75%  Appendix C  159  Experiment Worksheets  AUTOCLAVE TEST WORKSHEET  test ID: series ID: date:  #10 Copper Sunday, May 22, 1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 65 min 60 min  initial pressure: final pressure:  5.5 bar 5.5 bar  80 psig 80 psig  13.7 1/s  820 rpm  stirring rate:  1 h  REAGENT BALANCE Autoclave products  Autoclave feed IPC residue: electrolyte: distilled water:  leach residue:  360 g 700 mL 500 mL  leach filtrate: wash water: total filtrate: repulp volume:  299 g  residue ratio:  1018 mL 1142 mL 2160 mL  wash acidity:  1.20 wt./wt.  0 g/L H2S04  794 mL  P6TENTIAL M E A S U R E M E N T S before:  supernatant?:  106 mV (SCE)  351 mV (SHE)  after:  supernatant ?:  73 mV (SCE)  318 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 45 g/L  IPC residue of May 20,1994 55.35% 7.09% 9.13% 0.76% 5.04% 6.23% 14.72%  leach residue 76.68% 0.83% 0.57% 0.49% 0.39% 2.55% 17.50%  220 g/L IPC residue of May 20, 1994 8.88 3.76 2.36  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot: O  leach filtrate 0.32 22.6 30.6 1.26 16.7  g/L g/L g/L g/L g/L  0.005 M 0.383 M 0.521 M 0.017 M 0.299 M  69.0 g/L  0.703 M  repulp filtrate 0.247 0.496 0.664 0.021 0.309  g/L g/L g/L g/L g/L  Og/L  leach residue 30.07 4.38 6.86  METALLURGICAL B A L A N C E  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 86.6  90.3% 94.8% 46.5% 93.6% 232.8  calc'd head grade 55.03% 7.08% 9.12% 0.76% 5.04%  calc'd head extraction 90.3% 94.8% 46.7% 93.6%  balance errors (w.r.t. head grades) relative absolute 0.32% 0.58% 0.14% 0.01% 0.01% 0.10% 0.41% 0.00% 0.00% 0.06%  Appendix C  Experiment Worksheets  AUTOCLAVE TEST  test ID: series ID: date:  160  WORKSHEET  #9 Copper Saturday, May 21, 1994 320 °F  temperature: heat-up time to T: leaching time at T:  160 °C 64 min 60 min  initial pressure: final pressure:  5.5 bar 5.5 bar  80 psig 80 psig  13.7 1/s  820 rpm  stirring rate:  1h  REAGENT BALANCE Autoclave products  Autoclave feed 360 g 700 mL 500 mL  IPC residue: electrolyte: distilled water:  leach residue: leach filtrate: wash water: total filtrate: repulp volume:  POTENTIAL  301 g  residue ratio:  1012 mL 928 mL 1940 mL  wash acidity:  1.20 wt./wt.  0 g/L H2S04  516 mL  MEASUREMENTS  before:  supernatant ?:  105 mV (SCE)  350 mV (SHE)  after:  supernatant ?:  107 mV (SCE)  352 mV (SHE)  ASSAYS  component Cu Co Ni As Fe O Stot H2S04  electrolyte 60 g/L  IPC residue of May 20,1994 55.35% 7.09% 9.13% 0.76% 5.04% 6.23% 14.72%  leach residue 77.20% 0.82% 0.56% 0.43% 0.33% 2.13% 17.77%  220 g/L IPC residue of May 20,1994 8.88 3.76 2.36  element ratio (wt./wt.) Cu : 0 Cu : Stot Stot:O  leach filtrate 8.18 21.8 29.6 1.41 16.0  g/L g/L g/L g/L g/L  0.129 M 0.369 M 0.504 M 0.019 M 0.287 M  73.8 g/L  0.752 M  repulp filtrate 0.126 0.083 0.108 0.002 0.024  g/L g/L g/L g/L g/L  0 g/L  leach residue 36.24 4.34 8.34  METALLURGICAL BALANCE  component Cu Co Ni As Fe H2S04  assay head extraction  consumption kg/t IPC residue 93.7  90.3% 94.9% 52.7% 94.5% 220.3  calc'd head grade 55.18% 6.81% 8.78% 0.76% 4.77%  calc'd head extraction 89.9% 94.7% 52.4% 94.2%  balance errors (w.r.t. head grades) absolute relative 0.17% 0.30% 0.28% 4.00% 0.35% 3.78% 0.00% 0.53% 0.27% 5.27%  Appendix D  Air Discharge Calculations  161  APPENDIX D  AIR DISCHARGE CALCULATIONS I N C O ' s preliminary calculations [62] have shown that, in order to discharge a 6000 U S gallon First-Stage Leach ( F S L ) batch in 45 minutes, an air flow o f 107 scfm at 115 psig (8 kg/cm ) and 2  320°F would be required. A t the same flow rate it would take roughly 7 minutes to prepressurize the autoclave from 80 psig (5.5 kg/cm ) to the air discharge pressure. 2  The molar volume at 8 bar and 160°C is:  V  M  = 0.0224 x 8  x — 273  = 4.44xl0'  3  m /mole  (D-l)  3  1  From Appendix B it follows that a typical F S L batch has a volume o f 23150 L . K n o w i n g that the F S L autoclaves are operated at about 15% freeboard, the volume o f the plenum above such a batch can be calculated according to: 23150 Vptenm = -JJf - 23150 = 4085 L  (D-2)  Consequently, the total number o f moles o f gas present i n the reactor freeboard at the beginning o f the air discharge process would amount to:  N = — '° 4.44xlQ4  8  .  5  = 920 mole  (D-3)  3  The partial air pressure within the plenum at the initiation o f the discharge operation would be 2.5 bar, so that:  n  air  = — 8  x 920 = 287.5 mole  Approximating the composition o f air as 2 1 % (vol.) 0 "oxygen = 0.21  2  (D-4)  and 79% (vol.) N gives:  x 287.5 = 60.4 mole  2  (D-5)  Appendix D Air Discharge Calculations  162  Consider the oxidation reaction: Cu S 2  + H S0 2  4  + j 0  2  -> CuS0  4  + CuS +  (2-20)  HQ 2  The stoichiometry o f the above reaction dictates that two moles o f C u S 0 pass into solution for 4  every mole o f 0  2  that reacts. Furthermore, reaction (2-20) is kinetically fast; the cuprous sulfide  slurry can be expected to react with any oxygen introduced to the reactor almost instantaneously, whether it is stirred or not. Disregarding the hot 0  2  consumed during prepressurization, at least  120.8 moles o f cupric ion would be redissolved during air discharging o f a F S L autoclave. Hence, the minimal increase in copper concentration measured over the total volume o f electrolyte and steam condensate (Appendix B ) would be:  A[Cu ] 2+  =  120.8 x 63.5 21855  = 0.35  g/L  (D-6)  The actual increase in the copper content o f the F S L solution would even be higher, since the discharge pressure would be maintained at 8 bar.  

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