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By T. McCallum, B. Wellbrook, T. Ratnaweera

Leaching of copper (Cu) concentrates produces elevated concentrations of Cu and acid, which makes solvent extraction (SX) processing more challenging. Maintaining extraction targets can be more difficult with higher pregnant leach solution (PLS) grades, and variability of PLS Cu and acid can lead to increased chemical iron transfer. Diluting concentrate leach feeds with an existing heap leach inventory makes SX processing easier. However, considerations about how much to dilute and where to distribute the flow into the SX configuration must be carefully evaluated, given their impact on metallurgical performance. In this paper, characteristics of a concentrate leach feed are discussed, along with considerations for processing by SX. Also reviewed are pilot plant results of concentrate leach distribution into SX and plans for commercial testing. INTRODUCTION Hydrometallurgical processing of copper concentrates continues to gain interest from the industry. Treatment by smelting and electrorefining continues to dominate, accounting for ~80% of global copper production, but is faced with its own challenges (Li et al, 2017). The case for increased hydrometallurgical processing has been discussed as high grade ore deposits are diminishing, capital costs for smelting are high, and the environmental aspects present challenges for sustainable practice. The hydromet copper industry is also turning a corner as easily leachable copper oxide deposits are depleting, and excess SX-EW capacity exists in certain regions. The transition to hydrometallurgical processing of copper concentrates will take a bit of effort as leaching of copper sulfides is difficult and byproduct credits must be considered. Various POX leaching processes have been studied and proposed to the industry as viable practices. Of the many processes, commercial use has been practiced with the Mount Gordon Copper Process, total and partial pressure oxidation by Freeport-McMoRan at their Bagdad and Morenci, Arizona complexes, the Sepon Copper Process in Laos, and the bioleach BIOCOP™ process in Chile (Dreisinger, 2016).

Jan 1, 2019

By J. Y. Kim, Kenneth Stowe, Stephen Chryssoulis

"A series of flotation tests was carried out on Brunswick Mining ore with various pH modifiers (soda ash, lime, KOH) and added ions (ferric, nitrate, sulphate) in the copper lead rougher circuit. The main objective was to identify downstream effects on sphalerite activation and subsequent zinc rougher flotation. Some significant differences in zinc flotation were obtained. Most notably, soda ash and nitrate ions had a positive impact and ferric ions a pronounced negative effect on selectivity.Selected products from the flotation tests were deslimed at 400 mesh, washed with methanol, air dried, placed in teflon vials purged with nitrogen and stored at -15 degrees C. Numerous liberated sphalerite and pyrite particles were selected under a stereoscope and placed in rows on indium foil. The surface compositions of the particles were analyzed with an innovative laser probe (LIMS). The very small analysis area of the LIMS enables the surfaces of individual grains to be characterized. In addition, the rapid analysis time allows a statistically significant number of grains to be analyzed quickly.Surface lead appears to play a major role in zinc activation/flotation. Possible mechanisms are proposed to explain the variations in metallurgical performance at the different test conditions.IntroductionAlthough the bulk composition of mineral grains is an important consideration for the mineral processor it is ultimately surface composition and topography (edges and dislocations preferred sites for reactions) that control reactivity in flotation and cyanidation (passivation, encapsulation and preg-robbing). Determination of the surface properties of particles from process streams poses a particular challenge for the surface chemist for a variety of reasons, including:"

Jan 1, 1994

By Brien A. Perry, Art Wiebert

"Many organizations which have implemented Total Quality Management (TQM) are finding that the process has either stalled or is not working so well as originally envisioned. This can be particularly evident in an engineering environment where the uniqueness of engineers and their particular culture has not been adequately considered. Winning the commitment of a highly educated group ojprojemnnJJ1s to the TQM concepts, which are often initially 5W2 as too simplistic poses special challenges.Successful TQM implementations are found among technical groups where employees have significant input into creating the new work culture, feel empowered to make changes and are well trained in teamwork. A coach who is knowledgeable in implementation problems can, when working closely with the technical teams, help them circumvent obstacles and guide them successfully through the difficult early phases. The remaining key ingredient for success is supportive management which leads the change and is both informed and appreciative of the accomplishment being made.This Paper is based on first-hand experience with common problems and results encountered in implementing TQM in an engineering environment. Described in the paper are ways of avoiding these problems through awareness of the unique nature of engineering groups, implementation planning, just-in-time training and coaching people in the TQM process. The paper concludes with a focus on the opportunities and results from a successful implementation. These opportunities and results include the benefits each engineer can realize through process thinking, empowerment and teamwork."

Jan 1, 1995

By G. A. Sutton

This paper, describes a case example from San Manuel mine, where in situ copper leaching of the supergene zone was done over a 14 year period (1986–1999) in conjunction with underground and open-pit mining operations of a copper porphyry deposit. In 1993–1994, testing of well fields was done to determine the amount of copper that would be recovered from in situ leaching of the supergene zone, which was composed primarily of chrysocolla and copper-clay minerals. It was determined that the leach fluid would reach 70% (‘sweep efficiency’) of the copper minerals and that 75% of them would be leached for a recovery of 52% of the acid soluble copper. However, the final results from Zone 6, which had the longest leach history, showed that approximately 42% of the acid soluble copper was recovered, indicating that the `sweep efficiency’ was closer to 55%. Comparison of drill-core assay data from two holes, one drilled before (ARM158) and the other after (SMO9607) leaching, also indicate that approximately 42% of the acid soluble copper was recovered. The discrepancy between the predicted (52%) and actual (42%) recovery is probably related to a combination of factors, including difficulty in maintaining fluid saturation, channeling effects and gypsum precipitation. INTRODUCTION In situ copper leaching at the San Manuel mine, Arizona, was initiated in 1986. It was recognized that some of the copper oxide resource was not going to be economically recoverable by conventional open pit and heap leach methods using solvent extraction and electrowinning (SXEW). A high stripping ratio would be required to access part of the resource. At the time, in situ leaching was becoming popular in the uranium industry but the approach was virtually unknown in the copper industry, especially not at a large scale.

Jan 1, 2019

By E. J. Marcotte

"BRIEF DESCRIPTION OF GRINDING CIRCUITCrushed ore at -5/ 16 in. is delivered by a conveyor system from the crushing plant to two 1250 ton fine ore bins. The ore is transferred from each bin by a 30 in. feeder conveyor belt, operated by a Link Belt P.I. V. (variable speed) drive, to a 24 in. conveyor that feeds the 8 x 10 Allis Chalmers rod mill.The rod mill discharges to a common mill discharge pump box. Three 6 x 6 SRL pumps are employed to pump the combined mill discharge pulp to a common distributor box which distributes the feed to two of the three Kreb D lSB cyclones in each of the two banks at a hydraulic pressure of 6 psi.The cyclone overflow at a specific gravity of 1. 25 (i.e. 32% solids) is transferred by gravity to two revolving chip screens for wood chip removal and from there to a pump box. Two 5 x 5 SRL pumps feed the pulp to the thickeners. The underflow from the cyclones is fed to two 8 x 8 Allis Chalmers ball mills which discharge to the common mill discharge pump box. Thus, a closed circuit is formed between the ball mills and their respective bank of cyclones. The circulating load as most recently calculated is 262%."

Jan 1, 1966

By J. C. O'Rourke

"Gibraltar Mines Ltd. has been operating its open pit copper mine for fifteen years. In that time over 140 million tons of waste and low grade material has been placed on the mine waste dumps. Laboratory and in-situ testing indicated that enough leachable copper was recoverable from the dumps to justify a leaching operation which would include the construction of a solvent extraction-electrowinning plant. The plant, which was commissioned in October, 1984 was designed to process 3,000 usgpm of dilute copper solution to produce 5,000 tons per year of high quality cathode copper.This paper deals with the start up and initial year of operation of Gibraltar's SX-EW plant. The three areas: leaching solvent extraction and electrowinning are presented with emphasis on the innovations which were made to help overcome the problems unique to die Gibraltar operation.THE LEACHING CIRCUITThe mines waste dumps are leached to produce a pregnant copper solution as feed to the SX-EW plant. To achieve this the dumps are sprinkled with a mildly acidic solution which percolates through the dump. In travelling through the dumps the solution contacts the surfaces of the liberated copper minerals and disolves the soluble copper into the form of copper sulphate. The leaching reaction is achieved by a combination of bacterial action to oxidize the suphides and solublization of the copper into the acid solution.For the bacterial action to occur the dumps must be at or near the ideal pH of 1.9 - 2.1 and the necessary elements of the reaction must be present. Because a bacterial reaction is involved there is a lag time between the time when the conditions for bacterial growth are achieved and when the bacteria are in their active growth phase. Laboratory tests with Gibraltar ore indicated that bacterial action takes two to three months to reach its maximum rate. See figure 1.The expected lag time for the planned rate of bacterial leaching prompted Gibraltar to start irrigating the dumps in late June 1986 for a plant start up in early fall. The lag time proved to be longer than expected and the dumps took until November to start producing copper at the expected rate. See figure 2."

Jan 1, 1988

By J. L. Jambor, M. J. Baker, M. J. Logsdon

"Abstract - Prior to the development of the Diavik Diamonds Project, baseline studies were conducted to determine the geochemical characteristics of four kimberlite orebodies as an aid in the design of both the water-management system and the facilities for containment of processed kimberlite and ore stockpiles. Materials tested included field samples of volcaniclastic and pyroclastic kimberlite, processed kimberlite (i.e., kimberlite ore which had been screened and washed as part of the processing procedure), and sedimentary mudstone (a minor xenolithic unit which was assimilated during kimberlite emplacement). Approximately 200 samples of kimberlite materials were collected as part of the geochemistry program. Test-work included whole-rock chemical analyses, acidbase accounting, kinetic leach tests using columns, and mineralogical analyses. Diavik kimberlite has major oxide and trace-element concentrations consistent with global averages for kimberlite. The mean total-sulfur content of the kimberlite material is 0.22 wt% S, but with a significant range. The kimberlite has an excess of carbonate minerals over sulfide minerals (average CO2 = 4 wt%, present mainly as calcite), and has a mean neutralization potential of 311 kg CaCO3 equivalent/tonne. A reactive form of framboidal pyrite associated specifically with the mudstone xenoliths is the primary source of sulfide-sulfur. Long-term kinetic tests confirmed the preliminary interpretations that were made from the static-test results. Kimberlite and processed kimberlite are net acid-consuming materials that produce alkaline drainages and have low but detectable leaching rates for SO4 and specific trace metals such as Al, Co, Cu, Ni, and Zn. If segregated from the kimberlite, mudstone xenoliths are acid-generating (pH = 3) and produce an effluent with elevated concentrations of SO4, Fe, Al, Cu, Ni, and Zn. The study demonstrates that xenolithic units in Diavik kimberlites have an important effect on the environmental geochemistry of the ore rock. The mineralogy and aqueous geochemistry of the kimberlite materials are such that they may not be suitable for general earthworks or as an alkaline agent and should report to an engineered facility to protect site water quality."

Jan 1, 2001

By Rick Steen, Weldon Read, Morrism Tracy T.

Edging for starter sheet blanks used in conventional copper refineries has always been problematic for getting the edging to seal and adhere to the titanium material used for the blank so that copper is prevented from plating under the edging and cause undo problems in getting the copper sheets to separate from the blanks. A method was developed at the Amarillo Copper Refinery utilizing an injection molding process that better adheres the edging material to the titanium that makes the harvesting of starter sheets much easier to the point that robots are now used to strip with and which holds up very well to the unforgiving physical nature of the electroplating process. It is believed that this technology could be successfully utilized for permanent cathodes as well. INTRODUCTION ASARCO LLC (Asarco) was organized in 1899 as the American Smelting and Refining Company. Originally a consolidation of several lead-silver smelting companies, it evolved over the years into an integrated producer of primary non-ferrous metals, including copper, lead and zinc and associated co- products (Marcosson, 1949). Grupo México, S.A.B. de C.V acquired Asarco in 1999. Today, Asarco is a vertically integrated producer of primary refined copper and associated co-products such as gold, silver, molybdenum, selenium and tellurium. Asarco operates three open-pit mines and mills and two solvent extraction-electrowinning (SX-EW) plants all in Arizona plus a copper smelter in Arizona and a copper refinery in Texas. Figure 1 illustrates how the copper metal comes out of the ground at the mine and ends up as a refined product at the refinery. The copper refinery in Texas is known as the Amarillo Copper Refinery (ACR) and is located in Amarillo, Texas. It was built in 1975 to replace 3 other aging refineries and was designed to produce 550,000 tons (500,000 MT) of refined copper per year using conventional starter sheet technology. At the time it was commissioned it was the largest tankhouse in the world at just under 1/2 miles long under one roof.

Jan 1, 2019

By H. E. (Buzz) Neal

"Abstract - The Labrador Trough contains world-class iron deposits which have been mined since 1954. The direct shipping of Knob Lake ores were mined at Schefferville from 1954 to 1982. Concentrate production began in 1961 in the southwest end of the Trough, with three mines currently producing concentrate and pellets at a rate of 35 Mt per year. West and north of Schefferville, several billion tonnes of taconite have been outlined in fine-grained, cherty magnetite iron formation. Several deposits of highly metamorphosed magnetite-specularite iron formation are located west of Ungava Bay. Numerous studies have shown that this medium- to fine-grained iron formation can be beneficated to 66% to 68% iron. In the area from Wabush Lake to Mont-Wright, a medium- to coarse-grained friable specularite-quartz iron formation is repeated by folding to form several large deposits. In this area, three mining operations are located with reserves and resources in the order of 5 billion tonnes.The iron deposits occur within iron formation in sediments in the extensive Proterozoic geosyncline called the Labrador Trough. Several facies of iron formation within the Sokoman Formation reflect variations in chemical composition and depositional conditions. The correlation between the mineralogy and textures is discussed to illustrate the development of different types of iron ores. This band extends for about 1100 km southeast of Ungava Bay through both Quebec and Labrador. Further south, it turns southwest past the Wabush and Mont-Wright areas to within 300 km of the St. Lawrence River. The iron formation is essentially folded and faulted along most of its length. The degree of metamorphism is variable, ranging from intense in the northern and southern portions to greenschist facies in the central portion.The stratigraphy, structure and mineralogy of the producing mines are described along with the history of discoveries. Geological input is required to achieve the strict specifications of the international iron ore market."

Jan 1, 2000

Mining companies around the globe have long struggled with the need to ensure that earth moving equipment is designed to be operated and maintained under all site conditions without causing harm to people. Incident and accident data from a range of sources clearly identify mining equipment related issues as major causal factors. One Australian source, the NSW Department of Primary Industries? International Mining Fatality Database Project Report (2008) noted that ?Approximately 77% of mining accidents resulting in a fatality have? involved equipment?. The largest number of fatalities were associated with trucks; the next with LHDs. Many of the technical advances in equipment design over the past few decades have not been matched by advances in the human factors design of mining equipment, resulting in a ?design vacuum? that must be filled prior to commissioning new equipment. This ?design vacuum? is generally filled by local dealers. They retrofit equipment according to customer requirements to mitigate risks that have not been designed out by the Original Equipment Manufacturer (OEM). Mining companies have been frustrated by lost time and additional costs associated with this need to retrofit ex-factory designs. Some companies have attempted to address the problem by negotiating company or site-wide ?solutions? to equipment designs, but this has not simplified or clarified the issue from the point of view of the equipment manufacturers. A number of mining companies operating in Australia first discussed the concept of a joint customer approach to improve the Human Factors design of earth moving equipment at the factory level in 2004. They agreed that a new approach to engaging with OEMs was needed; one that would define the ?Problem?, in other words, the basis of the risks, rather than dictate the ?Solution?. The Earth Moving Equipment Safety Round Table (EMESRT) was formally established in 2006 by six global mining companies to establish a process of engagement between Original Equipment Manufacturers (OEMs) and mining customers.

May 1, 2009

By P. Mattar

The geotechnical importance of studying fluid flow through porous media and soils in particular, becomes evident when dealing with groundwater movement in surficial geomaterials that can be initiated by hydraulic gradients that develop during construction activity, groundwater extraction, impounding of reservoirs and groundwater recharge during adverse climatic events. In a geotechnical context, rocks are regarded as virtually impermeable materials, with comparatively greater resistance to flow of water through their pore structure. The importance of the permeability of rocks comes to the forefront when deep geologic formations are used in geoenvironmental activities for disposal of hazardous and toxic materials. The permeability of the rocks encountered is of critical importance to these geoenvironmental activities since the movement or transport of the hazardous materials during their eventual release will be largely governed by the permeability characteristics of the geologic formation. The estimation of the permeability characteristic of geologic formations is a non-trivial exercise since the measure of permeability can be influenced by the scale at which the measurement is made. Regional-scale and borehole-scale estimates of the permeability can be greatly influenced by the inhomogeneities, including fractures, fissures, solution channels and other anomalies that are present in naturally occurring geologic media. These anomalies are considered to be site specific and can only be determined in a reliable fashion by either conducting tests at the appropriate scale or in the regional-scale through inverse analysis of tracer movement patterns in rock masses. The estimation of the permeability of intact units of a geologic medium has important geoenvironmental applications in connection with storage of wastes since the units in which repositories will be located should be sparsely fractured and the sealing of the fractures and fissures are a prerequisite to developing an efficient solution for storage. The intact permeability of geomaterials is therefore regarded as an important property that influences efficiency and reliability of geoenvironmental solutions for deep geologic disposal. The permeability of geomaterials such as concrete and cementitious grouts are also important to many geoenvironmental and infrastructure applications where durability of structural materials in general and cements used for sealing deep boreholes and other access shafts employed in waste management endeavours is influenced by the permeability characteristics of these materials

May 1, 2009

By Simon C. Dominy, Ian M. Platten

"Abstract - Gold-bearing quartz reefs commonly show extremely erratic and unpredictable grade variation, although gross geological continuity may be good. This type of variation is often described as being nuggety or having a high nugget effect and can be measured quantitatively using the semivariogram. Understanding of geological features such as reef texture and structure will provide improved models for the interpretation of assay data, drill core descriptions, etc. In this paper, a case study from the quartz-hosted gold occurrences in the Gwynfynydd mine, United Kingdom, is described. The work provides a starting point for models of lateral variation and demonstrates the types of structural and textural features that may be sought to give clues to the prediction of gold distribution in similar deposits. Gold deposition at Gwynfynydd is primarily associated with the reef footwall. The host veins form a discrete group of structures that may have been emplaced early in the sequence of quartz veins forming the reef. This potentially exposes them to a maximum number of dissection events during repeat dilations of the lode. Gold was deposited in the local porosity at extreme concentrations, equivalent to kg/t grades, when the host fissure had become filled with a porous framework of crystalline quartz and sulfides. Distribution of porosity varied laterally along the vein as the result of the formation of clump-like growths of minerals from the vein walls. The growth style is relevant to the distribution of gold although it may not be relevant to the occurrence of gold in a particular vein. The preservation of pores with some connectivity late in the vein fill sequence may be important in permitting continued but slow fluid flows. This facilitates effective fluid reaction with wall rock-derived methane, thus changing the relative rates of gold and quartz deposition in favor of gold. The textural studies explain the first-order control of nugget distribution at Gwynfynydd but do not allow for prediction and optimization of the resource estimation process. This is principally because of the very low geological continuity and poor predictability of the high-grade pockets. Nonetheless, the work does provide a clear geological explanation for the erratic nature of the gold. It indicates textures and structures that can be used to determine geological continuity of the goldbearing elements within the gross reef envelope."

Jan 1, 2001

By Wei Ye, Lei Li, Xianjian Guo, Feng He

The purpose of the paper is to study the smelting process of copper concentrate with high Cu/S ratio produced in Africa or other areas around the world and theoretically analyze the influence of slag type on the smelting process. The paper has compared the processes used successfully in smelters, including CAPEX and OPEX and other factors, to provide reference for the industry to select the best process under specific conditions. INTRODUCTION At present, a great amount of copper concentrates with high Cu/S ratio, with chalcocite being the major source of copper, have been produced or are about to be produced in the Copperbelt in Africa and other countries in the world. The copper concentrate is usually high in copper (~30–55wt.%) and low in sulfur (~10–15wt.%). The conventional processes of smelting-converting and roasting-leaching-electro-winning for treating this type of concentrate is high energy and materials consumption such as electric power\fuels\flux et al. The direct-to-blister (one-step) smelting process can be considered favorable due to a reduction in processing steps. Up to now, there are four copper smelters in the world operating with the direct-to-blister smelting process. There certainly are problems in the process such as high copper content in the slag, strict operation requirement of slag cleaning electric furnace, etc. Except technical issues, the impact of factors on economical aspect such as power and material cost needs to be considered comprehensively in process selection. In this paper the technical issues of direct-to-blister smelting process, including slag treatment process, were analyzed and the comparison to the conventional processes and economic analysis of direct- to-blister smelting process to treat high Cu/S ratio concentrate were done.

Jan 1, 2019

By David L. Huston, Wouter Bleeker, Bruce E. Taylor, Ron Cook, Bob Stewart, Elizabeth R. Koopman

"Abstract—Detailed oxygen isotope mapping of rhyolites in and around the Kidd Creek mine indicates that the orebodies are associated with relatively low (<12‰) whole rock values that extend up to 200 m laterally from ore. The zone of low d18O values is overlain stratigraphically by a carapace of higher (>13‰) values that extends up to 1.5 km laterally from and 300 m to 500 m above ore. Immediately above the stratigraphically highest orebody this carapace is cut by a zone of lower d18O values extending from the ore zone. Zones of higher d18O values also occur in massive rhyolite bodies that occur in the footwall of the orebodies.The zone of low d18O values partly coincides with a Na2O depletion anomaly associated with hydrothermal alteration, except that the Na2O depletion anomaly has a greater lateral extent, and the low d18O anomaly extends farther into the stratigraphic hanging wall. These data suggest that d18O mapping may detect hanging wall alteration above massive sulfide bodies more effectively than conventional lithogeochemistry. Zones of high d18O values, which define zones through which large quantities of low-temperature hydrothermal fluids have passed, may also have usefulness in exploration.Oxygen isotope variations at Kidd Creek record three stages of a hydrothermal system that evolved in temperature with time, as follows: (1) an early, low-temperature event that produced rocks with high d18O values, (2) a high-temperature event that produced rocks with lower d18O values, and (3) a later, low-temperature event that also produced rocks with high d18O values. The high d18O signature of the massive rhyolite was recorded during early-stage, low-temperature (<200°C) alteration, and was probably preserved owing to decreases in permeability associated with low-temperature silicification. The low d18O signature of cherty breccia and chloritite was recorded during the thermal peak at >300°C. High d18O values in the carapace above the deposit were recorded as the hydrothermal system waned."

Jan 1, 1995

By Daniel Hodouin, Mélanie Proulx, Claude Bazin

"Collector and frother are commonly used to induce the floatability of valuable minerals in flotation plants. These reagents are usually distributed down a flotation bank or sometimes added in a single location at the top of the bank. A review of the literature and some experimental results obtained from laboratory and full scale trials indicate that the way flotation reagents are distributed down a flotation bank may influence the flotation response of coarse particles. The paper summarizes these observations which show that the distribution of reagents is a potential tuning variable for process optimization.IntroductionFlotation is in many cases the only way to concentrate sulphide and some oxide minerals. Flotation is a concentration process that uses particle surface composition to induce the selectivity by mean of hydrophobic reagents. The influence of particle size on flotation was firstly analyzed in the 50's1·2 and intensively investigated in the 70's, early 80's and 90's3-18. Most of the observed results show the existence of an optimum flotation size range9 whose recovery is not significantly influenced by a reduction of collector dosage""18. This result brought the idea of using a small portion of the total collector dosage at the top of a rougher bank and to distribute the remaining through the end of the bank to ensure sufficient collection of the coarse particles that need more hydrophobic coverage to attach to air bubbles and be lifted to the surface.This paper discusses the idea of using the distribution of reagents down a flotation bank to tune the operation of the process. Unlike the literature on the effect of reagents (collectors, frother...) dosage that is dense, little attention has been put on the distribution of reagents in a flotation bank. Although research on this subject has not been pushed very far, some operators have recognized the potential of a strategic addition of reagents down a flotation bank19. Except for a few exceptions, the proportions of the reagents to be distributed down a flotation bank are decided according to operator's experience instead of process analysis in relationships with particle size. This paper recalls observations made by different researchers that lead to the idea of using the distribution of reagents as a tuning variable for the flotation process. The paper does not present a detailed analysis or optimization scheme for tuning the distribution of reagents, but analyzes some plant and laboratory results with a view toward the development of a testing methodology and ultimately a validation or rejection of the idea."

Jan 1, 1998

Heap leaching accounts for around 20% of the global production of copper. In heap leaching coarsely crushed ore particles are contacted with an aqueous lixiviant percolating through large packed beds (heaps) of the ore particles, with the loaded solution collected at the foot of the heap for metal recovery, solution make-up and recycle. While dominant for copper oxide acid leaching, it finds application also in copper sulphide (bio) leaching. Despite their apparent simplicity, heaps are in fact complex chemical reactors in which a number of competing sub-processes determine the overall rate of metal extraction. Therefore, meaningful prediction of copper production rates from heaps, as well as operational optimisation and control, all hinge on a comprehensive and quantitative understanding of these complexities. The paper gives a comprehensive overview of the various sub-processes in copper heap leaching and their interactions, including those relating to mineral-particle-solution-gas phase interactions at the micro-scale bulk transport and distribution of solution in unsaturated dense beds at the macro-scale. On the basis of a more fundamental understanding and modelling, the potential for optimized operation offers new opportunities for the extraction of value from both low-grade and complex polymetallic ores. INTRODUCTION Heap leaching is one of the percolation leaching technologies, which includes in situ leaching, dump leaching, heap leaching and vat leaching (Bartlett, 1998). The common phenomenon in all these technologies is the transport of leach solution through a fixed bed of ore particles, which interacts with the solid as it migrates along its path. The target minerals are released into solution and with its flow are carried out of the bed. The targeted dissolved metal species are recovered from solution through various hydrometallurgical techniques, especially solvent extraction (SX), before the barren solution is returned to the leach process. Heap leaching is practised in three main types of applications: sulphuric acid leaching of copper oxide ores, cyanide leaching of gold ores as well as the oxidative acid leaching of secondary copper sulphide ores, usually with the assistance of certain microorganisms and, more recently, in the presence of chlorides (CaCl2, sea water). While gold heap leaching dates back to the 1960s, large scale copper heap leaching commenced in Chile from the 1980s, and sulphide heap leaching emerged essentially as a consequence of acid leaching from mixed oxide/sulphide ores from the 1990s (Watling, 2006).

Jan 1, 2019

By David L. Huston, Wouter Bleeker, Bruce E. Taylor, Ron Cook, Bob Stewart, Elizabeth R. Koopman

"Abstract-Detailed oxygen isotope mapping of rhyolites in and around the Kidd Creek mine indicates that the orebodies are associated with relatively low (<12‰) whole rock values that extend up to 200 m laterally from ore. The zone of low d18O values is overlain stratigraphically by a carapace of higher (>13‰) values that extends up to 1.5 km laterally from and 300 m to 500 m above ore. Immediately above the stratigraphically highest orebody this carapace is cut by a zone of lower d18O values extending from the ore zone. Zones of higher d18O values also occur in massive rhyolite bodies that occur in the footwall of the orebodies.The zone of low d18O values partly coincides with a Na2O depletion anomaly associated with hydrothermal alteration, except that the Na2O depletion anomaly has a greater lateral extent, and the low d18O anomaly extends farther into the stratigraphic hanging wall. These data suggest that d18O mapping may detect hanging wall alteration above massive sulfide bodies more effectively than conventional lithogeochemistry. Zones of high d18O values, which define zones through which large quantities of low-temperature hydrothermal fluids have passed, may also have usefulness in exploration.Oxygen isotope variations at Kidd Creek record three stages of a hydrothermal system that evolved in temperature with time, as follows: (1) an early, low-temperature event that produced rocks with high d18O values, (2) a high-temperature event that produced rocks with lower d18O values, and (3) a later, low-temperature event that also produced rocks with high d18O values. The high d18O sig- nature of the massive rhyolite was recorded during early-stage, low-temperature (<200°C) alteration, and was probably preserved owing to decreases in permeability associated with low-temperature silicification. The low d18O signature of cherty breccia and chloritite was recorded during the thermal peak at >300°C. High d18O values in the carapace above the deposit were recorded as the hydrothermal system waned."

Jan 1, 1995

By Xuanxue Mo, Kaihui Yang

"The Laochang mine, located in the Sanjiang area of southwestern Yunnan Province of China, is a volcanogenic massive Pb-Zn-Cu sulfide deposit. It is hosted by a pyroclastic sequence of Early Carboniferous age, namely the lower member of the Yiliu Group, which was intruded by syenite and rhyolite porphyry. The volcanic suite is basaltic or basaltic-andesitic in composition, has alkaline affInities, and is associated with the initial rifting along the margin of the Baoshan-Shanstate continental microplate (Yang, 1989). The ores are typically Pb>Zn>Cu in composition, averaging 3.820/0 Pb, 3.51% Zn, from 0.01% to 1.85% Cu (0.107% av.), a high average content of Ag (113 glt) and recoverable amounts of Au, n, Gd, Ga, Ge, In. A massive sulfide zone, in which lenticular and stratiform orebodies are localized in a sulfide-<:arbonate-silica shale unit or sulfidic tuffIte horizon that usually overlies coarse-grained pyroclastic rocks, is underlain by a stringer sulfide zone developed in a subvolcanic intrusion or volcanic sequence. The sulfIde mineral assemblages, from the top to the bottom of the deposit, are: realgar + orpiment - galena + sphalerite + pyrite - chalcopyrite + pyrite + pyrrhotite + arsenopyrite. Oxide and sulfate minerals occur rarely in the massive ores. The massive ores are generally fme-grained and banded, but coarse-grained ore, colloform and framboidal texture, and growth-zoned pyrite and sphalerite are common. Sedimentary features such as graded bedding, brecciated and deformed ore clasts, and chaotically emplaced blocks, are well preserved, indicating that the massive orebodies have undergone slumping and displacement. Alteration is weak in the hanging-wall rocks, but strong in the stringer zone. Two distinctive alteration zones have been identifIed: the upper zone, characterized by an assemblage of quartz, carbonates, chlorite, albite, zeolite and sericite, and the lower zone containing diopside, tremolite, actinolite, gamet, clinozoisite, and epidote.The general features of the Laochang deposit are in accord with the descriptive model of VMS deposits (Lydon, 1984). However, it falls into neither the Zn-Cu group nor the Zn-Pb-Cu group chemically (Franklin et al., 1981). The depositional setting of the Laochang deposit is quite different from Cyprus or Besshi-type massive sulfide deposits. Laochang is hosted by alkaline lavas that were erupted in a submarine continental rift. Although some features of the Laochang deposit are similar to those of Kuroko-type deposits, there are remarkable differences in the alteration mineral assemblage, the proportion of ore-forming elements (lead enriched) and the amounts of oxides and sulfates. These suggest corresponding differences in ore fluid composition and the associated ore-forming environment. Thus, the authors suggest that the Laochang deposit may be a distinct type of VMS deposit."

Jan 1, 1993

By Gerardo Fuentes

Arsenic extraction from Chilean copper concentrates up to nearly 99% was achieved, using hydrothermal reactions. Arsenic appears in nature in more than 200 mineral species, being arsenopyrite (FeAsS), enargite (Cu3AsS4), and tennantite (Cu12As4S13) the most common species. In addition, arsenic is present in a large amount of copper concentrates, produced in northern Chile. Medical statistics show that in northern Chile, the amount of cancer to the lung, skin and bladder are the highest in the country. The latter situation, demonstrates the importance of studying metallurgical processes that do not generate atmospheric pollution. The hydrometallurgical processes such as hydrothermal treatment in autoclave reactors, do not produce gaseous emissions. On the other hand, the arsenic remains in solution from where it can be transformed in scorodite by widely known methods. Previous works about hydrothermal treatment of enargite have shown the feasibility of transforming the enargite of the copper concentrates into chalcocite or mineral species of the chalcocite family such as djurleite, digenite and for this reason enrich the concentrate in copper and pass the arsenic to solution. For this research, two Chilean copper concentrates were used, concentrate 1 with 5.5% arsenic and concentrate 3 with 1.8% arsenic. These concentrates were subjected to a complete chemical and mineralogical characterization. Later these concentrates were carry out to batch tests of hydrothermal treatment. Also, these concentrates were subjected to grinding to have different particle sizes. For the hydrothermal tests, a liter capacity Autoclave Engineer reactor was used. Characteristic of this reactor was the maximum temperature 350°C, 5000 PSI pressure, and 900 rpm agitation speed. Kinetic tests were carried out with the two concentrates maintaining the same solid: liquid relationship, the same agitation velocity and the variables studied were treatment temperature, time of reaction, concentration of copper in the solution and grain size of the concentrates. The conclusions of the study indicated arsenic extraction of the order of 93.98% from the concentrate 1 and arsenic extraction of the order of 98.74% from concentrate 3. After the hydrothermal treatment, the solids and solutions were characterized chemically, by electron microscopy and X ray diffraction. Also the solids were characterized by grainsize through a particle size analyzer (MICROTRAC). Results of batch treatment shows that arsenic can be release almost completely from real copper concentrates using this hydrothermal process and the next stage will be a scale up process in order obtain results for pilot plant design. INTRODUCTION Arsenic is an impurity that appears associated with copper minerals forming enargite, tennantite, tetrahedrite, and that due to the pyrometallurgical treatment either in conversion or roasting furnaces, it goes into the atmosphere together with the sulfur dioxide. A great quantity of literature can be found to eliminate arsenic or try to separate arsenic from copper. The methods used are mainly leaching processes with or without pressure (Padilla, Jerez & Ruiz, 2015) with different reagents like chloride media (Herreros, Fuentes, Quiroz & Viñals, 2003) or chloride media with sulphuric acid (Riveros & Dutrizac, 2008) or sodium hypochlorite (Mihajlovic, Strbac, Zivkovic, Kovacevic & Stehernik, 2007). It is known too that Chile is the first copper producer in the world and a great quantity of its minerals are in the north of the country, in Antofagasta.

Jan 1, 2019

By Damian Connelly

The option to leach iron from low grade copper concentrates could reduce transport costs and significantly improve the saleability of the concentrate. In addition, it reduces smelting costs and can remove undesirable impurities. The trend to produce cleaner and higher grade copper concentrates is very appealing to the smelters. A recent copper project came on line and there were three serious deficiencies which were not identified in the feasibility study. This came as a surprise and resulted in a number of projects to investigate ways of rectifying the problems in the shortest timeframe possible and using minimal new capital. The first serious issued not identified when commissioning was the lower than expected copper recovery, which was 70% compared to a predicted recovery of 80% to 87% in the feasibility study. This required investigating new flotation reagents and incorporating controlled potential sulphidisation. The second serious issue overlooked is the presence of significant amounts of talc minerals (insol) in the ore which are naturally floatable. This leads to lower concentrate grade and attempts to remove the talc resulted in lower copper recovery. Smelters have limits on the amount of talc in concentrate because it produces viscous slags and difficulties for the smelter. It also incurs higher transport costs. The third serious issue not identified in the feasibility study was the high arsenic reporting to the concentrate. Arsenic is a penalty element for smelters and at levels exceeding 0.3% the smelters may reject the concentrate. Arsenic is a problem environmentally for copper smelters and penalties apply when it is high. This paper describes the options for penalty element removal and suitability to incorporate into the modified flowsheet. It also highlights where the geometallurgy was not properly addressed and how these surprises could have been avoided. INTRODUCTION From 1970 to 1985, a copper company undertook metallurgical testwork on the Peruvian project copper ore. The ores were mined underground and treated over many years previously. Various other groups undertook studies and work until Aker Kvaerner undertook a feasibility study in 2006. Three ore composites were tested and grind-recovery relationships were developed. The feasibility study testwork identified that talc was a possible concern and will have an impact on copper concentrate grade. This was to be managed by blending. The current operating reality is that there is so much talc in the ore body it is not possible to blend it out. There was far more emphasis on molybdenum recovery. No mention was made of arsenic issues other than blending. There was a Concentrate Leach Study and a Molybdenum Hydrometallurgical Plant Study but these options did not proceed to a final flowsheet. A Scoping Study in 2006 looked at a Heap Leach project. The heap leaching gave very poor copper recovery on the sulphide ore. Heap Leaching was evaluated and tested for the high talc and high arsenic ore but recoveries were poor. The Aker Kvaerner feasibility study opted for a simple primary crush SAG/ball grind followed by flotation and sale of a concentrate. This is the current flowsheet utilised and the scale of the project is very large based on processing 117,200 tonnes of ore per day. A feasibility study does not guarantee success and the deviations from the feasibility study suggest there is a lot of technical work required to improve the current metallurgical performance.

Jan 1, 2019