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By M. S. Moats

"The effects of commercial copper electrorefining additives (glue and thiourea) on the electrochemical removal of bismuth are of importance in copper production. It was found that Bi(III) electrodeposition occurs readily from acidic sulfate solution via a diffusion-controlled quasi-reversible reduction. The diffusion coefficient of Bi(III) was measured and found to be less than reported values of copper in similar solutions. The presence of thiourea and/or glue was found to accelerate or de-polarize Bi(III) reduction on stainless steel as compared to the inhibition or polarization of Cu(II) reduction. The additives also inhibited the dissolution of the Bi-Cu co-deposit during cyclic voltammetry. During short term deposition studies, the additives increased the concentration of Bi in co-deposited Cu-Bi products from solutions with low concentrations of Cu (<2 g/L). Bi deposition was significantly suppressed by increasing the Cu(II) concentration in the solution. The formation of a copper layer on the stainless steel electrode appeared to significantly suppress the reduction of Bi(III). This study demonstrates that glue and thiourea depolarize the deposition of bismuth on stainless steel. Therefore, the additives could be used to improve Bi impurity removal during third stage liberation.INTRODUCTIONThe production of high purity copper has been steadily increasing due to the wide application of the metal and its alloys in electrical and electronic products, infrastructure construction, industrial machinery and equipment, general products and transportation (Moskalyk et al., 2003). Most high purity copper is manufactured from copper sulfide mineral concentrates and/or scrap via a pyrometallurgical process followed by electrorefining (Navarro et al., 1999). During the electrorefining process, cast anodes of 99-99.5% purity are electrochemically dissolved into an acidified copper sulfate solution while high purity copper (>99.99%) is plated on stainless steel blanks or copper starter sheets (Davenport et al., 2002). The control of cathode deposit growth (purity and morphology) while maintaining a high current efficiency is of significant importance for copper refining (Ilkhchi et al., 2007). At typically applied current densities for industrial copper electrorefining, mass and charge transfer play important roles in the quality of cathodic products (Hiskey et al., 1998). If the real current density at some surface points is substantially higher owing to uneven current distribution, then rough spots can and will grow faster leading to nodules (Suarez et al., 1992). Nodulated deposits tend to capture solution and anode impurities (slimes) resulting in unacceptable cathode quality (Afifi et al., 1987). Furthermore, these nodules can grow until they eventually contact the anode leading to short circuits and current efficiency reduction (Ilkhchi et al. 2007)."

Jan 1, 2016

By K R. Brentley

"Tshepong Mine, Harmony’s largest gold producing mine, is situated in the Free State portion of the Witwatersrand Basin, east of Odendaalsrus and north of Welkom, South Africa. The Basal reef is a narrow tabular conglomerate and is the main gold bearing ore body. The strata strikes mainly in a northsouth direction with an average dip of 22 degrees. Currently mining takes place at depths of between 1 690m and 2 150m below surface. The Basal reef unit consists of a conglomerate ranging between a few and 30cm in thickness. Some 80-90% of the gold occurs in the lower portion of the conglomerate. Directly above the conglomerate is a quartzite placer unit, the Basal quartzite, which is a clean, hard competent rock and varies in thickness between 50cm and 200cm. The Basal unit is overlain by the Khaki shale which can vary between 30cm and 500cm in width. It is weak and incompetent khaki coloured chloritic shale.A method of stoping, termed “undercut”, is practiced in order to remove the lower 15-20cm of the gold rich strata. The method entails removing this portion along with some 70-80cm of footwall host rock which is a strong competent quartzite. The hard competent quartzite overlying the conglomerate is left in-situ to form a competent hangingwall which can stabilize and prevent exposure of the weak Khakishale. This method of extraction was developed through trial and error and has been used reasonably successfully since its inception in the early 1960’s where the method was developed to extract the Basal reef at the Western Holdings Mines. Recently prospect drilling and a number of localised hangingwall failures indicated that the residual quartzite beam could be less than 60cm and in some instances as thin as 30cm in certain areas of the mine. The result of this is a less stable hangingwall which not only poses a safety risk to workers in these areas but also had the potential of locking up some 840 000 Tonnes of gold bearing ore. At the present gold price this amounts to revenue of approximately ZAR1.5 billion (U.S. $143 million).This paper discusses the method of “undercut stoping” and existing practices in place at Tshepong Mine. It also explores new initiatives such as the use of a simple freely-supported elastic beam solution to perform back analysis, on-site data capturing of beam thickness and support geometries in order to develop empirical design charts and numerical modelling to assist in the verification of the empirical design method."

Jan 1, 2015

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

By I. Ueda, S. Ishikawa, T. Nagata

“O-SR process” in which the S-Furnace of the Mitsubishi process is followed by existing reverberatory furnaces (RFs) and Peirce Smith (PS) converters has been operated since December 2007. The S-Furnace smelts copper concentrates to produce molten matte containing 58-62%Cu which is then settled in the RFs and converted to blister copper in the PS converters. The SO2 gas generated from the S-Furnace and PS converters is converted to sulfuric acid in the acid plants. In addition to the smelting copper concentrates, the O-SR process has advantages in terms of recycling various waste/secondary materials. Automotive Shredder Residue (ASR) has been treated in the RF to reduce fossil-fuel consumption in RFs, and to recover various metals included in ASR as well, since 1993. Electronic circuit boards (E-scrap) has also been processed in order to recover not only copper but also rare resources such as gold, silver and PGMs since 2013. Strategic modifications have enabled expansion of the processing volume of E-scrap. Enhancing recycling of wastes and scraps from urban mines while reducing the consumption of natural resources in the “O-SR” process will be demonstrated and discussed in the present paper. INTRODUCTION Onahama smelter and refinery has been operating as a custom smelter since 1965. Smelting facilities were subsequently expanded to two green-charged RF furnaces and five PS converters in 1973. Since RF is a conventional old-fashioned smelting furnace that necessarily consumes a huge amount of fuel to smelt copper concentrates, saving fuel cost has been an important objective. In early 1980’s the fuel was changed from bunker oil to pulverized coal, while utilizing the combustion heat of waste tires. Furthermore, treatment of combustible waste such as shredder residue (SR) was initiated in order to save fossil fuel for smelting concentrates in the RFs. Since enhancing of the treatment of SR requires a reduction in concentrate smelted in the RFs, most concentrates were dried and injected through tuyeres into the PS.converters. These modifications enabled to reduce the amount of the pulverized coal combustion and to increase SR treatment in the RFs. SR treatment capacity has increased to 12,000mt/month because the SR market has grown after the enforcement of the Vehicles Recycling Law. Such business environment has been attributed to establish the thermal-materials recycling system of SR in which the RFs have changed their role from copper concentrates smelting furnaces to waste materials combustion furnaces.

Jan 1, 2019

By C. W. Chao, M. Melashvili, F. Hilário

The processing of a bulk concentrate containing copper as chalcocite, enargite, bornite and chalcopyrite along with pyrite and precious metals of silver and gold was studied. A leaching study was undertaken to examine the application of the atmospheric acid leaching process to a bulk copper concentrate for copper recovery followed by cyanidation or thiosulfate leaching for gold and silver recovery. The extraction of copper exceeded 90% under some conditions and the recovery of gold and silver reached a maximum of 44 and 45%, respectively. A sample of bulk concentrate was roasted in air to oxidize the sulfides with release of sulfur dioxide followed by acid leaching of the roaster calcine and cyanidation of the acid leach solution. The copper extraction reached over 92% and the gold extraction exceeded 88%. The sulfur dioxide may be captured for acid production. INTRODUCTION The processing of a large primary copper deposit in Indonesia is under investigation. One of the possible routes for processing involves flotation of a bulk copper-gold concentrate followed by treatment by an atmospheric leach/SX-EW process for copper recovery. The atmospheric leach residue would then be leached for recovery of gold and silver. A second route involves roasting of the concentrate followed by acid leaching and SX-EW for copper recovery. The leach residue from the roaster calcine would then be leached for recovery of gold and silver. The roaster off-gases could be used for sulfuric acid manufacture, enabling the production of a valuable by-product. The Cobre Las Cruces (CLC) process (Munoz, Garcia, & Delgado, 2017) was designed to treat a resource of copper discovered close to Seville, Spain. The process was developed by Outotec with MK Gold (original project owner). INMET Mining subsequently purchased the deposit from MK Gold and built the copper processing facility. The CLC process is similar to the Mt. Gordon Copper process; except that OKTOP reactors have been installed instead of a low temperature, low pressure autoclave process. The plant was designed to treat 1.3 mtpa of ore at a grade of 6.2% Cu to produce 72,000 tpa of cathode copper over a 15 year project life. The main gangue mineral at Cobre Las Cruces was pyrite. The Cobre Las Cruces ore is therefore similar to a rougher flotation concentrate containing copper minerals and pyrite. The roasting of a rougher flotation concentrate can be designed to produce hematite, basic copper oxides and sulfur dioxide off gas. The acid leaching of the basic copper oxides may produce soluble copper for SX-EW recovery of copper. The sulfur dioxide off-gas may be used to make sulfuric acid.

Jan 1, 2019

By Li Qihou, Liu Zhihong, Liu Zhiyong, Yao Wei

Black copper sludge, which mainly contains copper, arsenic and a little antimony, bismuth, selenium, tellurium, is an important recyclable secondary resource. In this paper, selective removal of arsenic from black copper sludge by alkaline oxidative pressure leaching was studied. The influences of alkali concentration, leaching time, leaching temperature, oxygen pressure, liquid-solid ratio and other variables on the leaching rate of arsenic, copper and tellurium were investigated. The results were as follows: when the leaching process was conducted at initial alkali concentration of 1mol/L, liquid-solid ratio of 10, stirring rate of 500r/min, leaching temperature of 140°C for 5 h and oxygen pressure of 2.5MPa, the leaching rates of arsenic and tellurium were 85.43% and 59.7% respectively, and copper, antimony, bismuth were not leached. Hence, arsenic has been removed effectively from black copper sludge, while the valuable metals such as copper, antimony and bismuth remained in the residue that can be further recovered easily. INTRODUCTION Black copper sludge is a byproduct in the purification process of copper electrolyte. Arsenic, antimony and bismuth are major harmful impurities in copper electrolytic refining system. The accumulation of these impurities will cause adverse effects, such as the increase of electrolyte resistance and power consumption and deterioration of cathode copper quality (Wang, Liu, & Zhang, 2001). Hence, electrolyte purification is an essential part of the electrolysis process (Liang, Chen, & Guo, 2009). Electrolyte purification methods include electrowinning (Davenport, King, Schlesinger, & Biswas, 2002), solvent extraction (Navarro, Simpson, & Alguacil, 1999), ion exchange (Arroyo et al., 2017), coprecipitation (Wang et al., 2003, 2018; Xiao et al., 2007; Xiao, Cao, Mao, Shen, & Ren, 2014), activated carbon adsorption (Navarro & Alguacil, 2002), electrodialysis (Cifuentes et al., 2015), redox process (Zhou, Peng, Zheng, Ma, & Cui, 2011; Peng, Zheng, Zhou, & Chen, 2012; Zheng, Peng, Ke, & Chen, 2013), among which the electrowinning method is simple in operation and effective in impurities removal, making it the main process for current electrolyte purification. Some copper ions can react with harmful impurities to form complex black copper sludge by controlling the operation conditions of electrowinning process. Based on the estimated data provide by Davenport et al. (2002), an annual output of 400,000 tons cathode copper could produce 432-2168 tons of black copper sludge by electrowinning method. Black copper sludge contains 30~50% copper, 15~30% arsenic and a small amount of rare elements such as selenium and tellurium. Thus, recovery of valuable metals can bring huge economic interests. Moreover, arsenic in these sludges could cause significant environmental pollution. Therefore, effective removal of arsenic from these sludges is urgent and significant.

Jan 1, 2019

By H. Fujihara, K. Uesugi, H. Toda, A. Takeuchi, K. Shimizu

"Hydrogen is trapped at a variety of microstructures, for example grain boundaries, dislocations and particles. It is thought that hydrogen trapped at trapping sites is redistributed to the other sites under loading and then brittle fracture occurs with hydrogen accumulation. In order to understand hydrogen embrittlement behavior, it is necessary to investigate the time evolution behavior of hydrogen accumulation at each trap site. In the present study, changes in hydrogen distribution are observed under stress using Kelvin force microscopy (KFM) that can visualize hydrogen concentration distribution on material surface at high resolution. Spot-like hydrogen enriched regions are observed on the surface of a material. In addition, hydrogen depleted regions are observed around those enriched regions, and the appearance of hydrogen trap due to particles existing near the material surface is revealed. On the other hand, hydrogen trapped at grain boundary can be visualized. From in this result, it was suggested that crystallographic factors influence hydrogen accumulation. The relationship between initiation and propagation of intergranular crack and hydrogen accumulation under stress will be discussed.INTRODUCTIONHydrogen present in a metal sometimes has significant effects on various properties. Hydrogen in a metal is trapped at a variety of microstructure, for example dislocations and grain boundaries. Grain boundaries, dislocations and micro pores are the major trap sites in aluminum alloys. Some researchers considered that hydrogen redistribution and accumulation between these trap sites in a material with loading plays an important role in the brittle fracture behavior. In general, in intergranular fracture and quasi-cleavage fracture, the hydrogen trapped at grain boundaries and dislocations contributes to fracture behavior, respectively. Recently some researchers have shown that the ductile fracture in aluminum alloys is resulted from growth and coalescence of hydrogen micro pores. Analysis using first principles calculation indicates that intergranular fracture occurs when hydrogen concentration reaches 20 atom H/nm2ongrain boundary. From this report, it is considered that hydrogen redistributed during deformation accumulates locally in order to cause hydrogen brittle fracture. In order to understand hydrogen embrittlement, it is necessary to investigate the time evolution behavior of hydrogen accumulation at each trap site. In the present study, we evaluate the relationship between hydrogen accumulation behavior and fracture behavior in Al-Zn-Mg based alloys. In-situ observation was made by combining the tensile test with Kelvin force microscope (KFM) that can visualize the hydrogen concentration distribution on the material surface with high resolution. We also investigate the relationship between hydrogen distribution behavior and crystal orientation."

Jan 1, 2018

By Mory M. Ghomshei

"The cost of deep underground mine cooling depends on a number of geological, environmental and economic factors such as the temperature gradient in the mine, the outside air temperature, and the local price of energy. While geological factors, being static, can be controlled to some extent during the mine design, the environmental and economic factors dynamically change and therefore can only be controlled through a flexible mine ventilation system.With an average temperature gradient of 20–25°C/km, the virgin rock temperature at 2,000 m can reach 45–60°C. Conventional cooling at such depths could lead to exorbitantly high costs. In-mine geothermal heat pumps can considerably reduce the cost of cooling through energy communication with the centralized absorption chillers, which service the mine from the surface.It is shown that in-mine heat pump systems can be a viable option for reducing operating expenses and providing some degree of flexibility to cope with fluctuating energy costs. The price of Natural gas and electrify are the most important variables for optimizing the level of interaction between the surface and inmine cooling systems. Deep mine waters can be used to carry a portion of the mine heat to the surface, where the heat from mine can be used as a supplementary energy source for absorption chillers. As an extra benefit, the heat output of the chillers can be used for district heating and or greenhouse growing.The case of cooling deep levels of deep mines in Sudbury Canada (e.g. Creighton Mine) has been discussed. It is shown that an initial 1MW in-mine geothermal heat pump system can save significantly on the cost of cooling. Such a demonstration project will allow R&D on viability of larger similar systems to be incrementally added in the future.INTRODUCTIONThe amount of heat generated in deep and ultra-deep mines depend primarily on the regional temperature gradient and the thermal conductivity of mine rocks. The conductive heat flow is quantifiable and can be minimized or optimized in the process of mine design. Another important heat source can be related to underground equipment such as diesel machinery which often have low efficiencies (approximately 25–35%). This heat source can be controlled and reduced though improving efficiencies or using electrical equipment, which offer a much better efficiency (approximately 80–90%). Mine waters can also be a significant heat source (convectively bringing large amounts of heat into the mine workings)."

Jan 1, 2019

By J. L. Jambor, E. J. Reardon, E. C. Appleyard, J. A. Cherry

Since the 1957 initiation of deposition of mill tailings at the Heath Steele mine, New Brunswick, periodic detailed studies of the tailings geochemistry and mineralogy have been conducted, thereby providing a unique opportunity to examine the geochemical evolution of a sulfide rich mine>tailings impoundment. When deposited, the tailings contained about 85 wt% sulfide minerals, 1 wt% to 3 wt% carbonate minerals, and 12 wt% to 14 wt% alumino silicate minerals. Mineralogical examination of samples of the tailings from deep in the impoundment indicates that little oxidation occurred prior to tailings deposition. The mill process-water deposited with the tailings was neutral to slightly basic and contained low concentrations of Cu, Pb, Zn and As. Sampling of the tailings five years after deposition indicated that sulfide-oxidation reactions, occurring near the tailings surface, had decreased the tailings-water pH and increased the concentrations of Fe, S042-, Cu, Zn and Pb (up to 10 000 mglL, 25 000 mg/L, 2.5 mg/L, 200 mg/L, and 4.5 mg/L, respectively). In 1975, ten years after decommissioning, the tailings could be divided into three zones: an upper oxidation zone in which sulfide oxidation is actively occurring, a hardpan layer, 15 cm to 20 cm thick, composed of tailings cemented by secondary Fe minerals, and an underlying reduction zone in which the tailings are relatively tin oxidized (Boorman and Watson, 1976). Chemical analyses indicated the presence of low-pH waters containing very high concentrations of SOi-, Fe, and other metals (up to 63 600 mg/L SO42-, 28 600 mg/L Fe, 10 mg/L Pb, 600 mg/L Cu and 4000 mg/L Zn), suggesting that there was a period of intense sulfide oxidation between five and ten years after tailings deposition. More recent sampling, conducted some 20 years after the end of tailings deposition, indicates that relatively little change has occurred in the last ten years, thereby suggesting that the hardpan layer has provided an impediment to the downward movement of dissolved constituents. Analyses of the tailings solids show increased accumulations of many metals at or near the depth of the hardpan layer. Evaluations of the long-term evolution of the tailings can be enhanced by comparison of descriptions of the tailings with descriptions of the gossan that overlay the massivesulfide orebody prior to mining. At Heath Steele, this comparison suggests that many of the mechanisms controlling the distribution of metals in the tailings will probably persist over geologic time.

Jan 1, 1992

By Louise Corriveau

Exploration strategies have traditionally considered felsic gneiss complexes of the Canadian Shield sterile. This paradigm is challenged with the discovery of extensive granulite-facies Cu-Au, Fe oxides-bearing hydrothermal systems in the 1.4 Ga Bondy and 1.5 Ga La Romaine volcano-plutonic complexes of the Grenville Province. Early identification of magnetite-rich units in association with meta-exhalite and pelite-like units was key to targeting the search for mineralization, protolith information and indicators of pre-metamorphic hydrothermal activity during regional surveys of these frontier terrains. Textural evidences for a volcano-plutonic setting were found. There remained, however, numerous impediments to the recognition that the ubiquitous magnetite-laden units and associated aluminous gneiss were hydrothermal in origin. Many metamorphic reactions form magnetite at the amphibolite to granulite facies transition, and hydrothermally-leached rocks resemble metapelite or migmatitic restite, while carbonate alteration zones may look like marble or calc-silicate rock. Non-typical metamorphic mineral assemblages, modes and colours arise when hydrothermally-derived rock types are metamorphosed. These effective markers served as tools for reliable field identification of hydrothermally-altered rocks. In our case-studies, we traced pre-metamorphic hydrothermally-driven Fe-oxide precipitation in association with sericitization, albitization, pyrophyllite alteration, chloritization, and B enrichment by mapping unusually high contents of (1) magnetite or hematite, (2) biotite, (3) albite, (4) sillimanite, (5) orthopyroxene or phlogopite with cordierite, and (6) tourmaline or kornerupine. The unveiled garnetite, tourmalinite, and aluminous gneiss had assemblages that vary according to metamorphic grade and composition. To assess if high mode of Fe-Mg-Al phases resulted from partial melting (restitic residue) or metamorphism of hydrothermal rocks, we compared leucosome mineral assemblages, shape and amount with that expected of in situ partial melting reactions. In both areas there is no compelling evidence for these units to be restites. In contrast, the presence of lapilli in garnet-cordierite-sillimanite gneiss, though extremely rare, constitutes conclusive evidence for extreme pre-metamorphic high-sulphidation alteration of a volcanic protolith, while Fe-oxides or sillimanite-bearing nodules and veins record a less pervasive alteration. Geochemical research on element mobility and mass balance calculations further testified of an hydrothermal origin for these gneisses. The magnetite-rich units, together with calc-silicate rocks, garnetite, hornblendite, albitite among amphibolite, provided effective guides to sites of fluid discharge and led to the discovery of Cu and Cu-Au showings. With the presence of highsulphidation alteration, these features suggest early Cu-Au VHMS-type alteration halos within a predominantly felsic composite volcano-plutonic arc setting for both case-studies. Evidences for subsequent F-driven orthomagmatic fluid activity are also present.

May 1, 2003

By W. K. Muwais, R. G. Cuddy

"The Lower Cretaceous McMurray Formation contains the bitumen reserves of the Syncrude Mine in northeastern Alberta. This formation comprises sediments deposited sequentially in fluvial, estuarine and marine depositional environments. The Syncrude Base Mine area consists mostly of estuarine sediments, as much of the original marine sediments have been eroded away. These estuarine deposits consist of channel-fill sediments and interchannel tidal flat sediments with the channel-fills comprising most of the orebody.Channel-fills contain, among other lithologies, interbedded sand, silt and clay rich layers that often occur as sequences of inclined beds with dips up to 25 degrees or more. A detailed knowledge of the distribution of these inclined beds, their lithological characteristics and their continuity is very important in assessing the risk of mine highwall instability . Techniques used to identify and characterize adverse structures are reviewed. INTRODUCTIONGeneralThe Syncrude Canada Ltd. oil sand mine is located within the surface mineable portion of the Athabasca Oil Sands area about 40 kilometres north of Fort McMurray in northeastern Alberta, Canada (Fig. I). The company commenced large-scale strip mining of oil sands in June 1977in the area, referred to as the Base Mine, situated in the southeast corner of the company&apos;s two lease areas (Fig. 2).The mining method (Fig. 3) utilizes large electric drag lines operating on a single bench with highwall depths up to 60 metres. Feed grade oil sand (= 6 wt.% bitumen) is mined by draglines and windrowed parallel to the highwall crest. Lower grade oil sand « 6 wt. 0J0 bitumen) occurring as waste bands is selectively mined by the draglines and cast back into the mined out pit or placed on the bench and hauled by truck. The windrowed feed is delivered to the extraction plant by a bucket wheel reclaimer and conveyor system. Stripping of Recent, Pleistocene and Cretaceous overburden units is done outside the conveyor systems by mobile equipment. The 4.5 kilometre long mine is split into four quadrants for planning and scheduling purposes, with each quadrant having an individual dragline-bucketwheel-conveyor system.Surface mining of oil sand at Syncrude requires the removal of vast quantites of material. The design of the open pit mine relies heavily on knowledge of the distribution of ore and waste materials, and on the physical characteristics of these materials.The geology of the oil sand below the dragline bench at the Syncrude Mine is complex. Geological, geotechnical and hydrogeological factors play an important role in the design and operation of the mine. This paper describes aspects of the geology that impact on the design and operation of the mine. Some of the methods and procedures adopted by Syncrude to successfully overcome difficulties that the complex geology creates are discussed."

Jan 1, 1989

By P. Espinoza, I. Yutronic, J. Olivares

"Compañía Minera Doña Inés de Collahuasi&apos;s (CMDIC) Concentrator Plant is located at 4300 meters above sea level, approximately 2000 kms north from Santiago, Chile. The concentrator has three grinding lines. Each one, Lines 1 and 2, have one SAG mill at the primary grinding stage followed by a secondary grinding and classification stage consisting of one ball mill, two hydrocyclone batteries and a sump with two pumps. Line 3 has one SAG mill (Gearless Mill Drive), followed by a secondary grinding and classification consisting two ball mills, four hydrocyclone batteries and a sump with four pumps. Average throughput for these three lines is 150 Ktons per day. Peak values of 180 ktons per day have been achieved.Since late 2008 and throughout 2009 and 2010, Operations and Automation areas developed an optimization program for the grinding process, considering the implementation of Model Predictive Controllers (MPC) for the three SAG mills, and later, extending MPC to the secondary grinding and classification stages. The implementation of a MPC solution was based on stability considerations (high variability of the SAG process, high interaction between sump level control and classification efficiency, etc), requirements for an automated and consistent operation, the need to maximize throughput under timevarying scenarios and, in the long term, optimization of grinding process considering Specific Energy Consumption, among other parameters. MPC solutions had proven control robustness and strong optimizing capabilities in terms of helping in discovering new operational points and so achieving all of the aforementioned goals.On the first stage of the program, MPC implementation considered identification of dynamical process control models based mainly on historical data, in order to know the impact of a group of Manipulated Variables over Controlled Variables. Also, the impact of measured Disturbance Variables was modeled in order to take into account the high impact and interaction of these variables over the processes. At the end of this stage, 4 controllers were successfully operating for the 3 SAG mills and the secondary grinding and classification process. In response to challenges made by the working group, the second stage of the development of control applications in the grinding process required to achieve a design and implementation of a solution capable of governing the milling process, in addition to stabilization and optimization of both processes (SAG milling and secondary grinding), to integrate from a point of view of process control and automation. Therefore, in the second stage, MPC performance was continuously monitored and optimized through advanced tuning, inclusion of more process models (with proper step tests), redefinition of specific optimization criteria and adjustment to new process needs.Finally, successful MPC implementation at Collahuasi has proved that fast development and deployment of process control solutions is possible and allows immediate process stability, higher throughput, specific energy consumption reduction, and successful product size control, among many others."

Jan 1, 2012

By Louise Corriveau

A Targeted Geoscience Initiative transect in the eastern Grenville Province reveals an extensive VHMS-style cupriferous hydrothermal system among rhyolitic pyroclastic rocks of the Wakeham Group at La Romaine, Québec. The volcanic rocks outcrop sporadically along a < 10 km-wide by about 100 km-long belt correlated with the south-eastern extension of the Wakeham Group. This belt occurs among voluminous, magnetite-bearing, granitic gneiss. At Lake Musquaro, 1.51 Ga metarhyolitic to dacitic lapillistones and flows, locally with aluminous nodules and veins, are intruded by 1.49 Ga feldspar porphyries. This volcanic package was deposited on typical Wakeham Group arenite (< 1.52 Ga) interbedded with nodular, muscovite- and fibrolite-bearing gneiss, and minor marble hosting finely layered garnetiferous hornblendite, garnetite and magnetite-bearing coticule. The latter are interpreted as meta-exhalite and with the associated albitite (sodic alteration), aluminous veins (hydrothermal leaching) and magnetite-rich granitoids, they are key elements in the search for concealed mineral deposits in this area. At La Romaine, to the south, 1.5 Ga lapillistones outcrop structurally above metasediments and below a composite amphibolite unit of uncertain origin (lava or sills) with magnetite-rich (BIF?) units. An extensive zone of aluminous gneiss, with lapilli textures, provide evidences for severe hydrothermal leaching of the volcanic rocks prior to metamorphism. An adjacent siliceous vein-textured unit likely represent a zone of intense silicification (74 wt.% SiO2), while aluminous, hematite to magnetite-bearing veins in adjacent felsic gneiss (tuff?) preserve evidences for the sequence of alteration events. Despite granulite-facies metamorphism, the main characteristics of the amphibolite unit persist along the entire belt. This competent unit likely served as a cap rock; the garnetite, diopsidite and hornblendite it hosts provide effective guides to the sites of fluid discharge and mineralization. A number of Cu showings, modest in size, were discovered in this unit. Their systematic occurrence, in a structural-stratigraphic sense, indicates that this setting is attractive for mineral exploration despite high-grade metamorphism and deformation. Late-stage veins laden with fluorite crosscut these alteration zones, and are themselves metamorphosed, indicating that hydrothermal activity was polyphased. SHRIMP U-Pb ages obtained on detrital zircons (1 arenite) and on igneous zircons (5 volcanic rocks, 1 porphyry and 6 granitoids), pervasive arc-type geochemical signatures and presence of in situ rafts of barium-rich garnetite in the granitoids indicate that sedimentation and volcanism occurred between 1.52 and 1.49 Ga coevally with hydrothermal activity and emplacement of a sub-volcanic 1.5 Ga granitic batholith with early, 1.53 Ga tonalite intrusion. Such a geological context can sustain large-scale hydrothermal cells and development of VHMS, Cu-Fe oxides or SEDEX deposits.

May 1, 2003

By C. Sui, J. E. Nesset, M. Li, J. Y. Kim, S. L. Chryssoulis

"A plant study of zinc rougher flotation at Brunswick Mining was conducted using lime, soda ash and a combination of both as the pH modifiers in order to establish the effects on sphalerite and pyrite kinetics and mineral selectivity. The plant sampling and analysis was augmented by surface analysis using TOF-LIMS, SEM and XPS as well as fundamental precipitation and surface-charge studies in the sphalerite Ca/C03/S04 system.The results revealed the important role of carbonate ions in cleaning the sphalerite surface for effective activation by CuS04 to occur, and the beneficial role of combined soda ash and lime in increasing pyrite and galena depression.1.0 IntroductionThe studies reported in this paper were initiated as part of an investigation into the causes of a particularly pronounced seasonal deterioration in metallurgical performance at the Brunswick Mining concentrator during the summer of 1995. The concentrator treats a complex base metal ore having a typical feed composition of 9.0% Zn, 3.5% Pb, 0.4% Cu and 100g/t Ag at a current throughput of 9,000 TPD, although at the time of the problem in 1995, the average production rate was closer to 10,500 TPD. Zinc, lead, copper and bulk (zinc-lead) concentrates are produced with overall metal recoveries of 86% zinc, 63% lead, 60% copper and 60% silver; results which are typical of the lower recoveries associated with fine-grained, highly complex VMS and SEDEX type deposits. The main economic minerals of interest are an iron-containing (7% Fe) sphalerite, galena and chalcopyrite with the main diluent (about 60% wt) being pyrite.Figure 1 shows a simplified flowsheet of the Brunswick milling circuit while the seasonal decrease in summer metallurgy in 1995 is illustrated in Figure 2. Of significance is the rapid drop in copper and lead recovery in June while there is an apparent delay until late August-September before the zinc deterioration reaches its lowest point. A more typical seasonal effect within the experience of the authors is the reverse of the Brunswick situation, namely an improvement in summer performance thought to be associated with improved differential flotation kinetics during warmer weather leading to better mineral selectivity. The historical Brunswick experience was particularly severe during the summer of 1995 with estimated incremental losses of $10 Million for this period. The potential of the very significant economic benefits to be gained from understanding and hopefully improving the annual drop in performance led to an investigation involving the minesite metallurgical group, the Noranda Technology Centre (NTC) as well as researchers in Professor Jim Finch's group at McGill and the surface analytical services and interpretive skills of AMTEL."

Jan 1, 1998

By Mariam Melashvili, Chris Fleming, Rene Jackman

"A study was undertaken to develop a flowsheet for treating a high grade silver/gold ore body by cyanidation, CCD and Merrill Crowe precipitation, with tailings detoxification to meet stringent limits for cyanide and thiocyanate concentrations in the final discharge. In one of the flowsheets examined, free cyanide in the final leach liquor was recycled to the mill where it was hoped that it might react beneficially with fresh ore to reduce the cyanide makeup by an equivalent amount. However, laboratory results suggested that most of the cyanide recycled to the mill reacted with iron forming a soluble ferrocyanide complex. This would have had a negative impact on cost, not only because the cyanide recycled to the mill could not be claimed as a credit, but also because the presence of ferrocyanide in the final tailings meant copper sulphate had to be added to the SO2/air cyanide detoxification process to precipitate the iron cyanide. In an alternative flowsheet that was examined, recycling of cyanide to the mill was abandoned, and this allowed the milled slurry to be pre-aerated in a single mixing tank before cyanidation. During alkaline pre-aeration, any ferrous ions that are generated in the mill are oxidized to ferric ions, which, unlike ferrous ions, are inert in cyanide solution. This significantly reduced cyanide consumption and lowered costs in the detoxification of the tailings. In this paper, the cost implication of these two circuits is compared.INTRODUCTIONMost of the gold that is produced by hydrometallurgical processes is recovered by cyanide leaching. The reason for that lies in the superior thermodynamic stability of the gold cyanide complex and the favorable economics this confers on the process (Fleming, 2014). However, the gold in most of the gold-bearing ores processed is a very minor component, and therefore, the leaching response is often dominated by that of the host minerals.Many gold ores treated by the cyanidation process contain iron minerals. Cyanide reacts readily with some iron minerals forming ferrocyanide as one of the compounds. Ferrocyanide is stable when solutions are exposed to air and does not retard the dissolution of gold. In fact, it has been claimed that potassium ferrocyanide in the presence of oxygen is a weak solvent for gold (Hedley & Tabachnick, 1958). The main problem with the formation of this compound is the consumption of cyanide. Ferrocyanide contains six molecules of cyanide for each atom of iron {Fe(CN)64-} so the wasteful consumption of cyanide is greatly increased by its presence. Thiocyanate is also often produced during the reaction of certain iron sulphide minerals with cyanide, and contributes to the wasteful consumption of cyanide. In addition, environmental regulations in certain countries demand that thiocyanate be removed from the tailings, and this adds significantly to the cost of tailings treatment."

Jan 1, 2016

By Q. Shi, Q. Feng

"The influence mechanism of chlorite on the flotation of pyrrhotite was investigated through flotation tests, Zeta potential measurements, optical microscope photos, solution ion measurements and absorption measurements. It is observed that, the addition of chlorite influences the flotation of pyrrhotite seriously. The IEP (isoelectric point) of pyrrhotite and chlorite is 7.6 and 4.3 respectively, in the condition of slight alkalinity, both pyrrhotite and chlorite are negatively charged, and the hetero-aggregation will not occur. With the existence of oxygen in solution, the iron ions are dissolved from pyrrhotite surface because of the oxidation of pyrrhotite during pulp conditioning. The dissolved iron ions adsorb onto the surface of chlorite in the forms of iron hydroxyl and iron hydroxide, and the adsorption results in the opposite charge of chlorite surface. As a result, the hetero-aggregation between pyrrhotite and chlorite occurs. The adsorption of chlorite changes the surface property of pyrrhotite and restricts the adsorption of amyl xanthate onto pyrrhotite surface, and then decreases the flotability of pyrrhotite.INTRODUCTIONIn sulfide ores processing, magnesium silicate bearing gangue minerals often report to the concentrate causing downstream processing problems as well as increased smelting costs. Talc,chlorite and serpentine are gangue minerals commonly encountered in complex sulfide ores (Engel, Middlebrook, & Jameson, 1997; Pietrobon et al., 1997; Witney & Yan, 1997; Chen, Grano, Sobieraj, & Ralston, 1999a, 1999b; Bremmell, Fornasiero & Ralston, 2005; Senior & Thomas, 2005; Wang, Somasundaran, & Nagaraj, 2005; WieseHarris & Bradshaw, 2007). All of them are characterized with uniform crystal structures consisting of silicon-oxygen tetrahedron and magnesium-oxygen octahedron. The magnesium-oxygen octahedron layers are broken during the liberation of serpentine to expose the magnesium-oxygen bonds with strong ionicity and hydration, which results in natural hydrophilicity of serpentine. In water, the hydroxyl on serpentine surface ruptures and dissolves into the aqueous phase while the Mg2+ remains, resulting in surface positive charge (Tao, 2012). As serpentine has a positive potential at alkaline condition, it is therefore likely that it will attach through electrostatic attraction to the negatively charged mineral surface and reduce both the flotation rate and recovery of target minerals (Lu et al., 2011). For talc, characterized as good natural hydrophobicity, the liberation comes from the interlaminar fracture to remain molecular bonds on surface with weak hydration. Being naturally hydrophobic, talc easily reports to flotation concentrates, thus reducing concentrate grade (Shortridge, Harris, Bradshaw, & Koopal, 2000; Beattie, Huynh, Kaggwa, & Ralston, 2006). Meanwhile, both molecular bonds and magnesium-oxygen bonds are exposed on chlorite surface when liberation occurs, resulting in moderate natural hydrophobicity between serpentine and talc (Tao, 2012). And thus the mechanisms of influence of chlorite on sulfide ores are more difficult to study."

Jan 1, 2016

By B. McFadzean, I. Achaye, T. Chidzanira, J. Wiese

"Froth stability has an overriding effect on metallurgical recovery and grade for many ore types. Froth recoveries of less than 10% are common for low stability froths such as those in sparsely mineralized ores or in cells at the back end of flotation circuits. It is well known that, aside from chemical factors such as frothers and electrolytes, the mineral particles themselves have the dominant influence on froth stability. The particle properties of size and hydrophobicity are expected to have the greatest effects on froth stability. The aim of this study was to understand the effects of particle size on froth stability and to quantify these relationships.The study compared a synthetic ore of pyrite and quartz, with carefully controlled size ranges and hydrophobicity, with conventionally milled platinum-bearing UG2 ore and the bulk reverse flotation of hematite ore. A novel bench-scale continuous column flotation cell and continuous mechanical hybrid cell with large relative froth volumes, as well as a three-phase froth stability column were used to characterize the froths.It was found that froth stability could be related to the concentrate particle surface area by a Langmuir-type relationship. This is indicative that the mechanism by which particles stabilize the thin films in froths is similar to the mechanism by which surfactants lower surface tension. Different froth stabilities were found for different ores and this was thought to be the result of different particle sizes and distributions, which resulted in different packing densities in the thin films and Plateau borders. Interestingly, there was a clear relationship between the feed particle size and the froth stability even though froth stability was dependent on the concentrate particle surface area. It seems that the feed particle size is a good predictor of the surface area of particles in the concentrate. Concentrate particle size did not show a clear relationship with froth stability since it was already shown that it is both the size and the amount of the particles in the froth that contribute to the stability.INTRODUCTIONThe stability of the froth in froth flotation is important to the recovery and grade of the valuable mineral. Too stable a froth and the grade is reduced; not stable enough and the recovery may be compromised. Platinum-bearing ores with fragile, poorly mineralised froths are known to have froth recoveries that are routinely less than 10% (Crosbie et al, 2009). There are a number of chemical, physical and operating parameters that can affect froth stability, including reagent dosages, air flow rate, pulp density and particle properties. Arguably, the largest contribution to froth stability is from the particles that stabilise (or destabilise) the thin films in froths. Particle size, hydrophobicity and shape all affect froth stability to some extent."

Jan 1, 2016

By G. Vibert Douglas

THE prospecter and geologist in the field, in the search for gold, are very often faced with a real difficulty. The: mineralized rock _frequently contains gold which is so finely divided or associated with sulphides that it is not recognizable in the gold pan. In recent years, the practice has been to take samples, send them to an assay office, and sit down and wait for results. This paper describes a method of solving the difficulty of proving the presence of this fine gold. The investigation is still in an experimental stage; but certain results have been obtained which would seem to justify publication. We do not consider that the method outlined is quantitative, but we believe that qualitative results are assured. The procedure has the great merit of simplicity. A sample of approximately two pounds of the mineralized rock should be taken and powdered as fine as is practicable. This powder is theo panned and the heavy minerals concentrated. Approximately two grams of these concentrates are taken for the test (one gram can be roughly gauged in the field by heaping up the powder on a Canadian twenty-five cent piece). The material is placed in a test-tube and 6 c.c. of a 10 per cent solution of iodine in alcohol are added (one-third of a test-tube full is approximately equal to 6 c.c.). Heat gently for about ten minutes over a spirit lamp, but without actual boiling. Allow to stand undisturbed until the sediment has settled to the bottom. Remove the clear iodine solution with an eye dropper. Allow it to fall, drop after drop, on a warmed pyrex watch-glass (the drops should be of equal size), taking care that the successive drops fall in the same place and that each drop evaporates before the next drop falls. Heat is applied to the lower side of the watch-glass by means of a spirit lamp. If gold is present in amounts of $2.00 or more per ton, a gold mirror will be formed on the watch-glass. This process of dropping and evaporating should be continued until no clear solution remains. If no mirror has appeared when all the solution has been used, another 6 c.c. of the iodine solution should be added to the powder and the process repeated. The absence of a mirror after the double treatment denotes an amount of gold less than two dollars per ton at the present price, or that the ore contains no gold. The mirror obtained must be washed with nitric acid ( 1: 1) to remove any copper, silver, lead, zinc, and iron. The yellow mirror remaining is pure gold.

Jan 1, 1937

By Waghmarem Umesh V., Vinay Jain, Pradip, Beena Rai

"The hematite surface plays a vital role in several processes of industrial and technological importance, such as, in beneficiation of iron ores, catalysis, and geochemical processes. Therefore, it is imperative to have a clear description of the structure and composition of this surface, which forms the underlying basis for such a wide variety of complex phenomena. Particularly, from the perspective of beneficiation of Indian alumina-rich iron ore slimes and fines, a clear understanding of the surface structure and composition of hematite surface and its difference from the alumina surface will facilitate identification and/or design of selective reagents which can be used to separate the two minerals. In this work, the hydroxylated hematite (0001) surface was modeled using density functional theory (DFT). The computed planar relaxations are consistent with X-Ray crystal truncation rod diffraction results reported in the literature. It is demonstrated that only one type of surface, that is, the (OH)3-Fe-H3O3-R is stable under aqueous conditions, as the relaxations for the second type of termination: (OH)3-Fe-Fe-R do not match quantitatively with experiments suggesting that its coexistence with the (OH)3-Fe-H3O3-R surface is unlikely. A detailed analysis of this surface shows that two of the three (OH) groups are not connected to the Fe surface atom. Instead, they form H2O molecules by combining with the H-atoms on the O-layer beneath the Fe-terminated surface leaving the Fe surface-atom bonded to only one OH-group. This undercoordination of Fe-atoms on the surface explains its higher reactivity as compared to hydroxylated alumina.INTRODUCTIONThe hematite surface plays a vital role in several processes of industrial and technological importance. A few examples are: flocculation/froth flotation of iron ores (Ravishankar, Pradip & Khosla, 1995), catalysis (Muhler, Schlögl & Ertl, 1992; Samsonov, 1973), and geochemical processes (Brown, Henrich, Casey, Clark, Eggleston & Felmy et al., 1999; Sposito, 1989). Therefore, it is imperative to have a clear description of the structure and composition of the hematite surface, which form the underlying basis for such a wide variety of complex phenomena. Particularly, from the perspective of beneficiation of Indian alumina-rich iron ore slimes and fines, this is an important problem, as a clear understanding of the surface structure and composition of hematite surface and its difference from the alumina surface will facilitate identification and/or design of selective reagents which can be used to separate the two minerals. Our current research focus is at employing molecular modeling techniques to come up with such selective reagents and then test their efficacy in the practical scenario through selective dispersion – flocculation labscale experiments and subsequently plant trials. Obviously, the correct structure and composition of the hydroxylated hematite surface is a fundamental prerequisite for such investigations."

Jan 1, 2016

By R. Kong, X. W. Zhang

"High pressure (up to 50 MPa) isotropic compression tests were performed on marble, gneiss and sandstone specimens, of which represents the typical rocks, to investigate isotropic compression deformation characterization at room temperature. Experimental data with respect to volume change in loading and unloading are presented. An apparent volume change inflection point (sensitive stress) was observed at a lower pressure range for porous sandstone, indicating that the elastic-plastic behaviour arising in the isotropic compression stage, but for tight crystalline rock, the sensitive stress is much higher, not probed prior to 50 MPa. The irrecoverable volumetric strain in the unloading can be seen for sandstone as well. The anisotropic deformation was discussed in terms of the axial and radial strain measurement. It was found that essential features of mechanical anisotropy are the irreversible in deformation range. The deformation feature in the isotropic compression tests is a very complex issue due to the factors concerning the stress history, unloading in sampling and in-situ stress. INTRODUCTIONThe isotropic or hydrostatic compression condition that stress is compressive and has the same magnitude for all direction, which is expressed as s1=s2=s3 in principal stress space. A yield locus of soils subjected to the isotropic and anisotropic compression was widely known, for instance Cam Clay model, which forms a closure cap model. However, the similar isotropic compression yield curves for rocks, especially for hard rock, was not established due to the geo-stress in the rock mass is much less than the socalled isotropic yield stress. And hence, the previous experimental studies on the hard rock have been made tend to on the stress-strain-strength relationship (Handin & Hager, 1957; Edmond & Paterson, 1972; Menéndez et al., 1996). Perhaps the yield curves related to an increasing in the hydrostatic pressure is not forced to construct for hard rocks, undoubtedly, the deformation data in this processing is valuable to assess the bulk module (Fabre & Gustkiewicz, 1997), anisotropy (Niandou et al., 1997), field stress (Hakala et al., 2007). Furthermore, many rocks store much elastic volume strain energy link to the hydrostatic pressure exerted, following a deviator stress applying, to drive the unstable brittle fracture for hard rock. Latter was studied to interpret the incentive of rock burst. However, volume change characteristic in the isotropic compression behaviour have been scarcely studied experimentally to our knowledge, less application related to the fracture in hard rock as well. The mechanical behaviour in the isotropic compression of hard rock is notably affected by the original structure anisotropy, stress history and stress condition before sampling, all of which need reasonably assess. It is also a useful understanding to study isotropic compression that elastic and plastic volume change in the isotropic loading and unloading, and the micro crack evolution throughout it."

Jan 1, 2015