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By F. Rojas, M. Sanchez, J. Palacios, O. Pavez

Copper sulfide is the main mining resources in Chile, and most of this metal is obtained via pyrometallurgical extraction in seven copper smelter plants, which extends from the northern Atacama desert to the central part of the country. In 2002, 1,522,000 metric tons of fine copper were produced, with a notorious environmental impact due to the large amount of slag generated, including transportation and disposing problems. Therefore, there is a great interest in finding alternative uses in order to give additional value to the slag. This is a total project development concept, which consider residues as new resources where new components could be obtained, improving sustainability in mining activities and also ensuring clean production processes. Thus, the present work shows a review of copper slag characterization as well as several ways of utilization, including compact material for road and railway installation, part of admixtures in concrete.

Jan 1, 2004

By R. Brough

"Over 25 years ago, SKEGA pioneered the use of rubber linings in grinding mills and rubber is today widely accepted and almost standard material chosen for Regrind - and Secondary Hills. There are, however, only a few Primary Hills in operation with ""solid"" rubber linings because available materials and designs are often not cost efficient in comparison with steel alloys on a Cost per ton basis and/or downtime for liner replacement would be to a high.In other words, solid rubber has limits as any other material. However, 'SKEGA started to think, why not try to use the best properties of rubber as well as steel and combine the two materials together.Nearly 10 years ago SKECA started to test this concept in Rod Mills. A steel cap was vulcanized on to the leading face of the lifter bars, to be installed close to the feed end head (fig. 1)."

Jan 1, 1988

By Jochen Petersen, Diane Taggart

"At an average of 90.3% (for the period of 1 January 2009 to 30 September 2011), Zincor’s zinc recovery was well below the industry standard of 96%. The losses through the residues, in particular the iron residue, were the biggest contributors to the overall zinc loss. It was believed that this loss could be decreased with the implementation of seed recycling. A plant trial was run to test the implementation of a seed recycle. The rate of seed addition through recycle was increased over a 4-week period. The presence of seed decreased the insoluble zinc loss by providing additional surface area for iron to precipitate onto, and reducing coating of unleached neutralising agent. The soluble zinc loss was also decreased at low and moderate seed recycle volumes. The improvement was attributed to agglomeration growth of the recycled particles, which improves the dewatering characteristics of the precipitate. In contrast, the soluble zinc loss increased at high seed recycle rates. This result was attributed to particle crowding and a low wash ratio.INTRODUCTION The Zincor refinery, which closed in December 2011, was situated just outside of Springs in the Gauteng Province of South Africa and employed the conventional Roast-Leach-Electrowin process, which accounts for 75 to 80% of the annual primary zinc production in the world. The refinery fed up to 220 000 t/a of zinc sulphide concentrate. The major portion of the raw material was sourced in Southern Africa from the Rosh Pinah and Black Mountain mines. Between 20 000 and 40 000 t/a of material was imported to make up the deficit and to improve the grade of the feed material. The iron contained in the zinc concentrates, at an average of 6.03% (for the period of 1 January 2009 to 30 September 2011), resulted in the formation of zinc ferrite during roasting, which is insoluble under atmospheric leach conditions. At Zincor, most of the zinc ferrite was dissolved in the residue treatment plant at elevated temperature and high acid strength. The leach liquor from this aggressive leach contained between 15 to 30 g/L of iron, mostly in the ferric form, and was known as the hot iron solution (HIS). The presence of iron in the liquor had an adverse effect on the zinc operation, particularly on the current efficiency in the electrowinning circuit. Therefore the iron needed complete removal."

Jan 1, 2016

By Yves Breau, François Robichaud, Jacques McMullen, Pierre Pelletie

"The Expert System implemented at the East-Malartic mill controls a copper flotation circuit while taking into account the profitability of the operation. Treating a copper gold ore, the flow sheet integrates a flotation step followed by cyanide leach circuit. The Expert System calculates the actual net smelter revenue of the flotation circuit, accounts for the operating cost of the cyanidation circuit and defines optimal metallurgical operating targets to optimize the overall economics. To be realistic, probability factors and circuit diagnostic analysis are integrated in the operating target selection. The Expert System has proved to be successful, robust and flexible. It resulted in a copper concentrate grade increase of 1% while maintaining recovery. The use of the expert system also gives plant personnel good training procedures and promotes interaction with mill operator.INTRODUCTIONThe East-Malartic mill is located close to the town of Malartic in North Western Quebec, Canada. The mill was first commissioned in 1938. The flowsheet at that time was a straight cyanidation circuit. The original mill operated in that mode up to the discovery of the Bousquet 2 orebody in 1989 which consisted of a massive copper sulphide ore averaging 0.7% copper and 7 g/mt of gold. Bousquet 2 ore processing required a mill retrofit. Grinding capacity, gravity and copper flotation circuits were inserted prior to the cyanidation circuit in order to achieve 3,200 mtpd capacity at expected recoveries.In 1993, control of the flotation circuit was deemed critical in order to maintain the flotation performance and to reduce cyanide consumption since the copper sulphide contained in the Bousquet 2 ore is a high cyanide consumer (50% bornite). (McMullen, 1999) As a result, the implementation of advanced control strategies in order to increase the metallurgical profitability of the plant and to reduce the operating variability between operators was the most rational alternative. In 1994, Corem was chosen to perform the plant audit and to write the first set of rules. The on-line control strategy was started up in 1995. In 1996 and 1997 the Expert System was shutdown because of an extreme change in the ore coming from the milling of Bousquet #1 ore and the departure of the projects engineer. In 1998, the system was restarted and the results were impressive. Furthermore in May of 2001 a patent was issued for the control strategy. (Breau, 2001)"

Jan 1, 2002

By G. Couture, J. -F. Dubé, F. Letarte-Lavoie

"The Xstrata Nickel Raglan concentrator, located in Quebec’s Nunavik region, has been in operation since 1997. Originally designed to process 800,000 tonnes per year of high-grade nickel ore, the plant’s current capacity is 1.3 million tonnes per year due to various optimization projects. In 2011 and 2012, the Metallurgical group, with assistance from BBA, updated and optimized the variables affecting SAG throughput in order to optimize process stability while maintaining maximum plant throughput. This project consisted of reviewing the control strategy and completely updating the pre-existing Fuzzy Logic controller. This paper describes the successful update of the controller including a complete revision of control rules, a new process control strategy and the integration of new process variables, most notably mill rotational speed. It will also show how this new approach has provided a simple and robust way of delivering both an increase in throughput while reducing process variability.INTRODUCTIONThe Xstrata Nickel Raglan Mine concentrator is located in Northern Quebec along the 62th parallel. It has been in operation since 1997 and was originally designed to treat 800,000 tonnes of ore annually. The Raglan Agreement held with the Makivik Corporation and the Inuit communities limits the annual tonnage of ore that can be processed at 1.32 million tonnes.The ore is processed in a 24 foot variable speed, semi-autogenous grinding mill (SAG). The conversion from Autogenous Grinding (AG) to SAG was completed in 2006. Before the ore is processed by the SAG, its size is analyzed by a Wipfrag camera that provides a granulometry estimate. A cone crusher reduces the SAG’s discharge screen’s oversize material. A 14 foot diameter ball mill ensures final grinding before the flotation stage.Many control strategies were used at the SAG over the past few years. Fuzzy Logic was first used in 2007 in order to control tonnage feed to the SAG. This first version only measured the value of the variables without taking into account their rate of change. Revisions were made over the course of the last few years to improve controller performance. Other variables were added such as slurry density and crusher Closed Side Setting (CSS). The previous revision of the Fuzzy Logic was done in 2009 with a review of the rules only."

Jan 1, 2013

By Neri Roux, Chris Budgell, Allan Cramm, Leo Cheung

"Anaconda Mining Inc. is investigating potential improvements and expansions to their operating plant located on the Baie Verte Peninsula in Newfoundland. An extensive batch and continuous bench scale testing program was launched to investigate the efficiency of cyanide destruction under different conditions in order to improve tailings handling. The primary objective of the cyanide destruction study was to apply new technologies to develop a novel cyanide destruction system.Effluent and slurry from Anaconda Mining Inc.’s Point Rousse Project was received at RPC where test work was conducted. Various sodium metabisulfite (MBS) dosages were evaluated and the use of utilizing ozone and ultra violet (UV) light to oxidize cyanide was investigated. All the data was evaluated and mass balances were compiled to develop several flowsheets. Overall encouraging results were found using a combination of ozone and MBS.IntroductionAnaconda Mining Inc.’s Point Rousse project includes an operating gold processing plant located on the Baie Verte Peninsula in Newfoundland which is in the process of improving and expanding. This plant comprises of milling and flotation followed by cyanide leaching in order to extract the gold. The residue from cyanide leaching is filtered by drum filters prior to being clarified for the Merrill-Crowe process. Once the gold has been recovered by the Merrill-Crowe process the effluent is filtered by filter presses and the resulting filtrate as well as the solids from the drum filters are detoxified prior to being pumped to the tailings pond along with the flotation tailings.The entire process is very dependent upon the efficient destruction of cyanide in the effluents and successful tailings handling. Substantial amounts of effluent are delivered to tailings ponds after gold extraction which could create environmental problems if the regulations are not met. Environmental regulations stipulate that final effluents contain no more than 1 mg/L Total CN.The detoxified tailings of the Point Rousse project are produced by the INCO SO2-Air circuit depicted in Figure 1. These detox tailings contain 10 mg/L Total CN at that point of the circuit and this was selected as the target to be matched while reducing overall MBS consumption rates from 3.2 g/L per minute. Prior to detoxification, the filter press tailings contain 520 mg/L Total CN while the drum tailings contain 29 mg/L Total CN."

Jan 1, 2017

By B. S. Fee, R. R. Klimpel

"Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores such as cobalt, platinum, gold, silver, etc. Most of the initial developments made with the process of froth flotation in the early 20th century were based on the processing of metal sulfide ores. It is the relative technical and economic ease of floating metal sulfide minerals with small amounts of appropriate frothers and sulfur containing (thio) collectors that has been a major economic driving force behind the rapid expansion in the supply of metals such as copper, lead, zinc, and nickel that has occurred in the world's economy in the last 50 years. This author estimates that 1.6 billion tons of sulfide ores are processed each year (1980) in the free world based on the froth flotation process.While there has been a tendency for the quality of metal sulfide ores in any one geographic location to deteriorate over time due to mining intensity, the overall global supply of sulfide minerals is still very large. Thus, the physical supply side of sulfide ore concentrates has not been under strong technology improvement pressure with increasing metal demands. Part of this is due to the ability of capital to consistently move to new geographic locations having suitable grade sulfide ores. In these moves, the actual froth flotation technology used (including collector chemicals) has had to change little. As an illustration of this trend, of the approximately 80,000 metric tons of thio collectors used commercially (1980) in the free world, almost 98% were known and manufactured in some form 25 years ago (1). This is clearly not typical of reagent development in other process related industries. In addition, the industrial scale practice of froth flotation applied to sulfide ores has proceeded since the 1920 s with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully at an industrial scale for many years."

Jan 1, 1987

By Hans von Michaelis

"INTRODUCTIONThe renaissance in gold and silver recovery technology has continued to unfold and it is appropriate to review some of the useful new technologies as well as try to identify some of the industry's future needs and frontiers.The remarkable strengthening in the US dollar has resulted in a relatively low gold price in US Dollars of around $310/oz at the time of writing and silver at $6.25/oz. This general softening in the dollar prices and the recent confusion as to environmental compatibility of cyanidation tailings and effluents, has led to a slowdown in announcements of new operations. Another reason for the slackening off in the USA is that in the past few years a large part of the capital has come from oi I and gas companies that now seem to be losing interest in mineral ventures, and foreign investors for whom the dollar is becoming out of reach. In the author's opinion this represents an unprecedented window of opportunity.Elsewhere in the world where local currencies have weakened, the price of gold is at or close to its all time high.For example, in South African Rands, the gold price was R571.00/oz when the rand was US$1.47 and the gold price reached US$800. Now at US $309/oz and with the Rand at US$0.51 = Rl .00, the local gold price is R600. The price of si Iver in Mexican pesos is not much less than it was when silver reached US$50/oz. Gold, and silver prices in local currencies outside North America are still buovant and at levels where additional investment in new projects is more encouraging.Even in Canada where the dollar is worth about US$0. 75, this provides a smaller but significant cushion compared with US operations. Considering the present strength of the US dollar and even the Canadian dollar relative to many other gold producing countries, there should be significant investment opportunities at today's prices. Mining companies have started to descend on Brazil, for example, where in a matter of just a few years annual gold production has increased from 10 to 70 tons. In South Africa, a Canadian company backed by Swiss financing has been able to take over a Witwatersrand gold mining operation and is installing a huge new tailin9s retreatment plant to process about 1 million tons per month of tailings."

Jan 1, 1985

By Norman O. Lotter, Dominic Fragomeni, Lori J. Kormos, Guy Comeau, Antonio Di Feo

"A performance opportunity was identified for the Raglan concentrator operation by changing the selection of xanthate from Potassium Amyl Xanthate (PAX) to Potassium Isobutyl Xanthate (PIBX). This opportunity was identified by a Statistical Benchmark Survey of the operation and QEMSCAN measurement of these survey samples, in which fine liberated pentlandite was found in the final tailings. This was confirmed by some laboratory scale flotation testwork using High- Confidence Flotation Testing. A statistically designed plant trial was implemented in 2007 to measure and confirm the performance gain. This trial design used on-off switching following a three-month reference distribution of normal operations using PAX. This layout provided both external and internal reference distributions on PAX to be established. These were used to evaluate the internal reference distributions on PIBX. This trial format was deliberately designed so as to block out the effects of autocorrelation in the operations data. In this way, small but significant metal recovery gains may be measured and proven in a plant trial. The paper describes the case study to show that this approach successfully blocks out autocorrelation in time, which is an obstacle in analysing concentrator operations data. The results showed that the PIBX had produced clear and statistically significant recovery gains for Ni, Cu, Pt and Pd. Adjusted for feed and concentrate grades, these recovery gains were: Ni: 1.00, Cu: 0.96, Pt: 5.07 and Pd: 1.90%.INTRODUCTIONConcentrator OperationsMilling and flotation, sometimes accompanied by gravity separation, are well-established unit processes. These have traditionally been used to treat ore that has been mined, producing a lowbulk concentrate containing a high percentage of the paymetal(s) in the ore mined. These concentrates are more economically treated in subsequent beneficiation processes than the ore from which they were produced.This choice of a concentration process results in a significant and variable paymetal loss to the concentrator tailing. This tailing is discarded and disposed of in an impoundment area, and has no immediate business use. This paymetal loss forms the largest loss of paymetal in the entire flowsheet by which refined, or saleable, metal(s) are produced [Cramer, 2001]. The question arises as to whether any particular concentrator is performing at its full potential, and how this answer may be determined. In a concentrator operation, there is always an ongoing performance improvement programme for this reason."

Jan 1, 2009

By J. R. Goode

"The Las Cristinas deposit contains about 10.2 million ounces of gold at a grade of 1.3 g/t. The deposit comprises a layer of fully oxidized saprolite (SAPO), a layer of sulphide-enriched saprolite (SAPS), carbonate leached bedrock (CLB) and carbonate stable or un-leached bedrock (CSB). Gold occurs at about the same level in all lithological groups. The bedrock units contain minor amounts of copper (about 0.1%) as chalcopyrite, while SAPO contains virtually no copper it having been leached and deposited in the SAPS zone.A previous owner started investigation of the Las Cristinas deposit in Venezuela in 1991. Over the next seven years, 1174 holes with a total length of 159 km were drilled, extensive metallurgical testwork, including pilot plant operations, were conducted and feasibility studies were completed. Construction was started in 1997 and again in 1999 but was suspended.Crystallex has been producing gold in Venezuela since the early 1990’s. In 1997, Crystallex acquired Inversora Mael which had held two of the claims to the Las Cristinas concessions since 1986. In September 2002, Crystallex entered into a definitive agreement with the Corporación Venezolana de Guayana, (CVG), to develop the Las Cristinas deposit.An extensive program of studies and testwork was started in early 2003 to determine an optimum development plan for Las Cristinas. Mine Development Associates, SNC–Lavalin Engineers and Constructors, SGS Lakefield Research Limited, J.R. Goode and Associates, and Professor A. Laplante at McGill University acted as the main contractors.Because of the potential value of by-product copper, earlier flowsheets used carbon-in-leach for SAPO, but processed all other ore types by flotation to produce a gold-copper concentrate with cyanide leaching of cleaner tailings to maximize gold recovery. Cyanide recovery was necessitated by the high copper content of the leach feed. Initial studies of available data by Crystallex revealed that direct leaching of all ore types would provide about 11% more gold recovery than the previously selected flotation route with reduced capital and operating costs.Bench and pilot scale studies have confirmed that a SAG-ball mill-gravity-CIL route is very effective for all ore types and will give about 89% gold recovery. This paper describes the recent testwork, discusses the proposed plant design, and presents economic data."

Jan 1, 2004

By F. W. Brackebusch, R. O. Sack

"A method for determining the fineness of Au in electrum from the composition of fahlore coexisting with pyrite and chalcopyrite is presented. This method is based on previously published experimental brackets and activitycomposition models for electrum and fahlore, and it is potentially useful for defining temperatures of hydrothermal mineralization and retrograde evolution of precious metals deposits. The authors demonstrate that it successfully predicts compositions of electrum reported for this assemblage in the literature and illustrate its application for three gold-quartz vein deposits from the Coeur d’Alene mining district, Idaho, that are prospects of the New Jersey Mining Company, the Coleman vein, High-Grade prospect, and Silver Strand mine. Temperatures calculated from average electrum compositions estimated from gravity concentrates (molar Au/(Au+Ag) = 0.76-0.39) and from microprobe analyses of the fahlores (molar Ag/(Ag+Cu) = 0.039 ± 0.004, Zn/(Zn+Fe) = 0.57 ± 0.03, As/(As+Sb) = 0.23 ± 0.02) from the Coleman vein indicate that the electrums underwent retrograde re-equilibration to temperatures between about 80°C and 240°C, roughly 80°C below the temperature of hydrothermal mineralization and 10°C to 40°C above the temperature inferred for retrograde equilibration of Fe and Zn between sphalerite and fahlore. Fahlores from the High-Grade prospect [molar Ag/(Ag+Cu) = 0.009 ± 0.004, Zn/(Zn+Fe) = 0.55 ± 0.11, and As/(As+Sb) = 0.04 ± 0.03] and Silver Strand mineralized zone [molar Ag/(Ag+Cu) = 0.10 ± 0.04, Zn/(Zn+Fe) = 0.63 ± 0.02, and As/(As+Sb) = 0.06 ± 0.02] exhibit strong core to rim depletions in Ag due to retrograde operation of the reaction1/10Cu10Fe2Sb4S13 + Ag + FeS2 =Cu-fahlore electrum pyriteCuFeS2 + 1/10Ag10Fe2Sb4S13chalcopyrite Ag-fahlorewhich proceeds to the left with decreasing temperature. Electrum inclusions in arsenopyrite from the High-Grade prospect [molar Au/(Au+Ag) = 0.90 ± 0.02] define a minimum bound on the temperature of hydrothermal mineralization of 273°C, and the temperatureelectrum composition curve constructed for the most Ag-rich fahlore in the Silver Strand mine indicates that the fineness of its electrum is significantly less than in the other deposits."

Jan 1, 2004

By D. T. White

"Metallurgical extraction processes often call for the heating of aqueous slurries. Indirect heat exchangers are generally not feasible for service with high yield stress slurries because of low heat transfer rates. It is more appropriate to heat the slurry by direct contact, typically with steam as the heating medium. Although having a long and successful history, direct-contact heat transfer to aqueous slurries has been used without the support of pilot plant research. This paper provides results of the needed research, the work being done at a scale large enough to enable direct scale up to commercial operations. A pilot plant heater was constructed and tested to develop an understanding of the slurry heating process and to ensure that commercial scale heaters were appropriately sized.INTRODUCTIONPressure Leaching OperationsPressure leaching operations are often used to oxidise sulphide gold ores, or for extraction of nickel and cobalt from laterite ores with sulphuric acid. Where there is sufficient sulphur available, the exothermic pressure oxidation reactions can be self sustaining at temperatures as high as 225 ºC, without having to preheat the feed slurry. At lower sulphur levels, slurry preheating is required. This is typically achieved by utilising the flash steam released from the slurry letdown stages and contacting this directly with cold incoming slurry ahead of the autoclave. The vessels where this slurry-steam contact is made are known as heaters or condensers, with the number of pre-heat stages varying, depending on how much heat is evolved in the oxidation reaction (Mason, 1990).In nickel laterite pressure acid leaching processes, which operate between 245 and 270 ºC, the only significant exothermic reactions are those associated with dilution of sulphuric acid in the leach. As a consequence, slurry preheating with flash steam is always used commercially. In practice the number of flash steam preheating stages required varies from 1 to 3, driven by a number of factors, including other uses for flash steam and the upper temperature limit at which preheated slurry can be pumped. A diagram of a typical pressure acid leach train is shown in Figure 1.For pressure acid leaching operations and at low sulphur levels in pressure oxidation, boiler generated direct steam addition to the autoclave is required to sustain the reaction temperature. For this reason it is imperative that the flash steam – slurry heat transfer is efficient and the slurry temperature approaches that of the flash steam closely."

Jan 1, 2012

By D. R. Boyle

"The Murray Brook precious metal (Au, Ag) gossan deposit is one of nine supergene mineral deposits that have developed over polymetallic massive sulfide deposits in the Bathurst Camp of New Brunswick. Reserves consist of 1.9 Mt of gossan containing 1.53 g/t Au and 65.9 g/t Ag. The supergene zone at Murray Brook consists of six distinct units: a) altered massive sulfide, b) pyrite-quartz sand, c) massive sulfide gossan, d) disseminated sulfide gossan, e) ferruginized wallrock, and f) leached bedrock. The massive sulfide gossan constitutes the main body of economic mineralization and consists of goethite, primary quartz, secondary amorphous silica, K-Fe-Pb-As- Sb-Ag hydrated sulfate and oxide minerals (beudantite, plumbojarosite, jarosite, bindheimite, scorodite), trace cinnabar, and cassiterite of primary origin. The order of stability of sulfide minerals during oxidation was pyrite _ arsenopyrite _ galena _ Bi-Sb sulfosalts _ chalcopyrite _ tetrahedrite group _ sphalerite.Two mass balance techniques (isovolumetric and conservative element) indicate elemental depletion in gossan relative to primary ore in the order: S>Cd=Zn>Cu=Mn>Na=Ca>Fe=Al=V> Mg= K=Se=Hg>Pb>Ti>Ag>Mo=P>Ba; and enrichment in the order Au>As>Sb>Si>Bi. Tin, as cassiterite, is conservative.Mineralogic and paleomagnetic data suggest that the gossan deposit formed under a warm temperate climate throughout the Pliocene. The rate of oxidation and the compositional nature of the ore was significantly influenced by: a) the position of the primary sulfide zone in a downward moving hydrological system, b) primary sulfide composition and zonation that promoted the development of a strong electrochemical oxidation system, and c) pronounced intercalation of carbonate and sphalerite mineralization in sulfide zones, which promoted rapid development of secondary porosity, permeability and downward flow of oxygenated groundwater.During progressive oxidation and physico-chemical erosion of the gossan zone, Au was transported downward in groundwaters, probably as a Auo colloid complex, to be concentrated in the lower horizons of the gossan profile. The precipitation of Au was greatly increased in the pyrite/arsenopyrite zones. Gold concentration was multi-cyclic with continual dissolution and nucleation of the metal until final concentration in the void structures of the gossan. Leaching experiments and microprobe analyses indicate that Au is present in the gossan as submicron composite sols of Au-Ag-silica."

Jan 1, 1995

By D. R. Boyle

"Abstract - The Murray Brook precious metal (Au, Ag) gossan deposit is one of nine supergene mineral deposits that have developed over polymetallic massive sulfide deposits in the Bathurst Camp of New Brunswick. Reserves consist of 1.9 Mt of gossan containing 1.53 g/t Au and 65.9 g/t Ag. The supergene zone at Murray Brook consists of six distinct units: a) altered massive sulfide, b) pyrite-quartz sand, c) massive sulfide gossan, d) disseminated sulfide gossan, e) ferruginized wallrock, and f) leached bedrock. The massive sulfide gossan constitutes the main body of economic mineralization and consists of goethite, primary quartz, secondary amorphous silica, K-Fe-Pb-As- Sb-Ag hydrated sulfate and oxide minerals (beudantite, plumbojarosite, jarosite, bindheimite, scorodite), trace cinnabar, and cassiterite of primary origin. The order of stability of sulfide minerals during oxidation was pyrite > arsenopyrite > galena > Bi-Sb sulfosalts > chalcopyrite > tetrahedrite group > sphalerite.Two mass balance techniques (isovolumetric and conservative element) indicate elemental depletion in gossan relative to primary ore in the order: S>Cd=Zn>Cu=Mn>Na=Ca>Fe=Al=V> Mg= K=Se=Hg>Pb>Ti>Ag>Mo=P>Ba; and enrichment in the order Au>As>Sb>Si>Bi. Tin, as cassiterite, is conservative.Mineralogic and paleomagnetic data suggest that the gossan deposit formed under a warm temperate climate throughout the Pliocene. The rate of oxidation and the compositional nature of the ore was significantly influenced by: a) the position of the primary sulfide zone in a downward moving hydrological system, b) primary sulfide composition and zonation that promoted the development of a strong electrochemical oxidation system, and c) pronounced intercalation of carbonate and sphalerite mineralization in sulfide zones, which promoted rapid development of secondary porosity, permeability and downward flow of oxygenated groundwater.During progressive oxidation and physico-chemical erosion of the gossan zone, Au was trans- ported downward in groundwaters, probably as a Auo colloid complex, to be concentrated in the lower horizons of the gossan profile. The precipitation of Au was greatly increased in the pyrite/arsenopyrite zones. Gold concentration was multi-cyclic with continual dissolution and nucleation of the metal until final concentration in the void structures of the gossan. Leaching experiments and microprobe analyses indicate that Au is present in the gossan as submicron composite sols of Au-Ag-silica."

Jan 1, 1995

By P. Lind, A. Gagnon

"Goldcorp has embarked upon an ambitious, industry-leading, multi-year water use reduction effort. This is a proactive, voluntary effort that reflects our commitment to truly sustainable mining. This effort will focus on increasing water re-use and recycling, and reducing water inputs ultimately towards zero. All of our operating mines are expected to contribute to this effort and are currently in the process of developing plans to meet short and long term objectives. One of the largest needle movers to increase water re-use and recycling is in the area of tailings management. Within five to ten years at Goldcorp, it is envisioned that the need for wet tailings dams will be eliminated and that all new tailings facilities will consist of only filtered or dewatered tailings. Filtered tailings is developing into a mature technology with wide application in mine waste management. This paper describes a conceptual level evaluation on the transition from a slurry to a filtered tailings operation at the Cerro Negro mine in southern Argentina. Key benefits and challenges are highlighted.INTRODUCTION Goldcorp’s Cerro Negro operation is located in the Santa Cruz province in the Patagonian region of Argentina. Cerro Negro has been in operation since 2014 and commercial production was achieved on January 1, 2015. The processing capacity of the plant is 4,000 tonnes of ore per day and in 2016 produced 363,000 ounces of gold and 3.1 million ounces of silver (Goldcorp Inc., 2017). As the mine is located in a semi-arid area, it is important to evaluate technologies that could reduce freshwater consumption and maximize water re-use and recycle rates. The current tailings operation at Cerro Negro incorporates pumping of tailings slurry into a conventionally lined Tailings Storage Facility (TSF). Water is reclaimed from the tailings supernatant pond to the process plant via pumps, however, water losses occur due to evaporation and solution retention within the slurry tailings voids. Converting to a filtered or dewatered tailings operation would allow for increased water recovery prior to tailings disposal, and decreased water losses from evaporation and from retention in the tailings solids void space. This paper describes a conceptual level evaluation on the transition to filtered or dewatered tailings with the primary objective of moving closer Towards Zero Water."

Jan 1, 2017

By G. Q. Zhou, H. C. Liang, X. D. Zhao

"This study describes the strain behavior in vertical shaft lining (VSL) embedded in thick soil deposits and its application in assessment of concrete deterioration. The observed strain in VSL continuously decreased companied by a sinusoidal fluctuation due to the influences derived from negative skin friction (NSF) and temperature. The increase in strain increment between peak and trough at each year and compressive strain level as the servicing time continued and VSL depth increased implies that the concrete strength was attenuated, and this attenuation was significantly dependent on the existing compressive strain level in VSL. The reliability of the proposed approach was validated by Schmidt Hammer Technique (SHT) – based concrete strength, and is applicable in concrete assessment and then fracturing prediction of VLSs in thick soil deposits.Concrete strength, Field monitoring, Strain behavior, Vertical shaft liningINTRODUCTIONOver 100 VSLs in Southeastern China subjected to an unpredictable fracturing as shown in Figure1, and there are over 1,000 VSLs embedded in thick alluvium su?ering from challenges of damaging and potential fracturing (Cheng, Yang, Zhou, Li, & Hu, 1993; Zhou, Cui, Lv, Huang, & Chen, 1999; Li & Li,2005). Therefore, a reliable assessment of these VSLs is significant to the normal servicing of mine shafts and the recycle of underground space when these mine shafts were closed.The continuous increasing of NSF induced by the drainage of water–saturated deposits well interpreted such extensive fracturing phenomenon, and it is universally regarded as the governing indicator to determine the VSL stabilization (Cheng et al., 1993; Zhou et al., 1999). Based on this recognition, the fracturing prediction and VSL strengthening according to the field monitoring on hydrogeological characteristics, servicing environments, and mechanical behaviors of VSL were carried out (Zhao, Zhou, Zhong, Zhu, & Liang, 2009; Liang et al., 2009; Xie, Lv, & Wu, 2014). As known, the compressive strain level in VSL and the in–situ strength of VSL concrete were the main factors to perform such fracturing prediction. However, the existing studies in literatures were limited to the strain accumulating or releasing behaviors (Zhao, Zhou, Zhao, Kuang, & Hu, 2017), and lack of considering the VSL concrete strength and its attenuation. In this reported study, the in–situ monitoring data on the strain in typical VSLs and the VSL elevations between 2004 and 2014 were initially analyzed. Then, an observed strain behaviors–based approach was developed to quantitatively evaluate the VSL concrete deterioration."

Jan 1, 2019

By M. Nagamori, A. J. Plumpton

"A low-grade concentrate from the Crevier Township, Quebec, containing 1.8% Ta205 and 13.7% Nb205 was reduced with carbon at 1600°C, and niobium and tantalum were selectively absorbed in liquid iron. To slag off the zirconia-rich gangue, some silica and lime were added as flux. Metallic iron was added in such a way that the combined contents of TaC and NbC did not exceed 20% in the ferroalloy . Up to 95% of the niobium, and up to 80% of the tantalum contained in the concentrate were recovered in the metal. The ferroalloy was ground finely and dissolved in 2 wt % sulphuric acid to leach out iron and silicon preferentially. The insoluble residue consisted principally of NbC and TaC, the calcination of which yielded an enriched concentrate comprising 80% Nb2O5 and 8% Ta2O5. The insoluble residue could also be chlorinated to obtain an impure liquid of Nb2Cl5 and TaCl5. Process variables and their effects are discussed in terms of higher metal recoveries in the smelting.IntroductionPresently, about one-third of the world production of tantalum comes from the pyrometallurgical reprocessing of tin slag, a by-product of tin smelting, containing sometimes 2 to 10% Ta205 and about the same amount of Nb205. Upon heating tin slags with carbon at about 1600°C in an electric furnace, tantalum and niobium oxides are converted to TaC and NbC, which are absorbed by the coexisting phase of molten metallic iron(1,2). The Fe-Nb-Ta-C melts cannot be decarburized by oxidation because of the preferential formation of Nb and Ta oxides over CO gas(3). So, the ferroalloys containing tantalum and niobium carbides are commercially treated by acid-leaching and I or chlorination to eliminate iron and thus permitting the eventual recovery of purified tantalum and niobium oxides(2).Absorbing Ta and Nb carbides in molten iron was first proposed and demonstrated by Zimmerley and Back(4) in 1957 in order to recover selectively tantalum and niobium from lowgrade concentrates. The process was successfully applied to the electric furnace smelting of tin slags by Hunter and Fursmant(5) over a period preceding 1965."

Jan 1, 1985

By A. R. MacPherson

"This paper considers four aspects of autogenous grinding. Part 1 discusses statements on power efficiency of autogenous grinding compared to conventional crushing and grinding systems. It also demonstrates how high-power efficiency can be obtained from autogenous grinding systems. Part 2 considers a simple method of testing ores to determine the type of autogenous plant which can treat these ores on a commercial basis. A comparison of testing for simple small plants vs large plants or complex ores is included. Part 3 considers important aspects which have to be considered in proper design of an autogenous grinding system. Part 4 discusses problem ores and how they can be treated in an autogenous system. INTRODUCTIONAfter about 35 years of using autogenous grinding plants in the mining industry, it is perhaps time to review where we are with this, by now, relatively old development. Autogenous grinding in this paper means grinding run-of-mine ore by fully autogenous means or by using a ball charge to assist the grinding (semi-autogenous). Also, autogenous grinding can be eitherby dry or wet.The use of autogenous grinding in the fifties was making its slow start, and it was not until the sixties that more of the mining industry would consider its use . By the seventies, most builders of new plants would have considered autogenous grinding, whether they used it or not, in the final design.In the eighties, most large plants use autogenous mills and to reduce over-all operating costs, many older conventional plants are now looking and actually retrofitting these old plants with autogenous grinding mills.It is evident that many misconceptions concerning autogenous grinding are still current. Some come from improperly designed plants which result from poor test work. There are many considerations in the proper design of autogenous plants. Some ores have been found to give problems even in a well designed plant. These ore types are beginning to be classified and recognition in the test stage allows steps to be taken in the design so that these ores can be efficiently handled in autogenous grinding plants.Therefore, this paper will address the following aspects of autogenous grinding:1. Misconceptions regarding autogenous grinding.2. Testing of ores to ensure a well designed commercial plant.3. Important considerations in mill design.4. Difficult ores, their recognition and types of design to handle such ores."

Jan 1, 1989

By Takanori Kokubo, Kenta Suzuki, Tomohiro Sasaki, Toshiya Anami

"Previous work has shown that the strengthening Al-Mn alloys with copper leads to the formation of an undesirable reduction in recrystallized grain size. In this study the mechanisms of strengthening and recrystallization behavior in a continuously cast automotive heat exchanger fin stockof Al-1.1wt%Mn-0.5wt%Fe-0.3wt%Si with a0.3wt%Cu addition were investigated. Metallurgical analysis revealed that the Cu strengthening was dominated by the formation of dispersoid particles, while solid solution strengthening decreased with the Cu additions. This result is attributed to the decrease in the amount of Mn solid solution. The post brazing recrystallized grain size was reduced compared to that not containing Cu. It is postulated that the Cu addition leads to a decrease in the amount of Mn in solid solution which leads to a finer grain size. This effect is attributed to Mn in the aluminum matrix inhibiting boundary migration but allowing other grains to grow during heat treatment. The results are rationalized based on particle stimulated nucleation (PSN), wherein dispersoid particles are more likely to become active recrystallization nuclei during the heat treatment. The increase in the number of active dispersoid particles with Cu addition is correlated to the recrystallized grain structure.INTRODUCTIONImprovements in properties such as strength, braze ability and corrosion resistance in thermal fin stock alloys are all means to gain improved efficiency in heat exchangers particularly in automotive applications. While Al-Mn alloys dominate automotive heat exchanger fin stock alloy, even small chemical compositional changes can profoundly influence these properties (Kawaharaet al., 2004; Yoshino et al., 2009). The microstructure (particularly the grain size and shape) also plays a key role in determining the product performance and hence the thermo-mechanical processing (TMP) must also be considered. The casting method (typically conventional direct chill, DC, or continuous casting) significantly influences the properties in no small part due to the very different TMP routes that are used. Continuous casting is generally favored for fin stock applications as its higher cooling rate (which can be 2 to 3 orders of magnitude higher than DC casting) leads to finer as-solidified intermetallic particles and an increase in super-saturated solid solution elements (such as Mn)that are known to improve fin stock performance such as strength, braze ability(Okiet al., 2008)."

Jan 1, 2018

By G. Morin, F. Chiesa, B. Tougas, B. Duchesne, Jocelyn Baril, N. Giguère

"An aluminum A356 compressor impeller was sand cast using the Low Pressure technique. The casting was limited in size (170 mm in diameter, 1.2 kg in weight) and very intricate in shape thus making 3D mold printing particularly favorable; indeed, producing this part by the traditional pattern-core box route would involve mounting 12 cores inside the mold while stringent dimensional tolerances are required as the impeller blades are only 2 mm thick. Consequently the molds were produce on the X-One MFlex printer of Quebec Metallurgy Center - CMQ. A dozen castings were produced with different casting temperatures and filling times so that it was possible to assess under which conditions complete filling of the mold cavity would be achieved. Filling and solidification modeling allowed predicting misruns and metallurgical properties inside the A356 (AlSi7Mg04) casting.INTRODUCTION Recent advancements in additive manufacturing enable foundries to produce molds without the tooling (pattern and core boxes) necessary when conventional technology is used (Eyad, 2016). This is particularly advantageous when dealing with intricate shapes and/or when a limited number of parts must be produced within a few days rather than in 5 weeks. An additional advantage is gained when producing prototypes (Johnson, 2016) as changes are much less costly and time consuming when made on a numerical 3D drawing rather than on patterns and core boxes. 3D printing is normally economical for a small number of parts; however, if the part is not too big and very intricate in shape such as the part that will be studied in the present work, 3D printing can be economical even for an important number of castings to be produced. On the other hand, low pressure casting is a high integrity casting process in which the mold cavity is filled from the bottom via a transfer tube in which the liquid aluminum rises under the pressure applied at the surface of the melt as illustrated in Figure 1. The filling can thus be precisely controlled, eliminating the turbulence created when the mold is filled from the top (gravity casting). Also, the pressure applied on the melt during solidification results in excellent feeding of the casting when conditions of directional solidification are met. In the present work, the compressor impeller shown in Figure 2 will be sand cast by low pressure."

Jan 1, 2018