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By David deVries, James Rajith, Paul Mascrenhas, Nurzhan Dyussekenov, Lawrence Hanschar

Kansanshi Copper Smelter commenced operation in March 2015. The smelter features four Peirce- Smith Converters (PSCs) operated in a three hot, two blowing configuration. This paper describes some of the process automation and design features implemented to maximize operational and environmental performance. To ensure that the converters were easy to operate, efficient, and were protected in emergency situations such as power failures or loss of tuyere air supply a significant focus was placed on automated sequences initiated by “one button” operator commands or otherwise automatically by the Distributed Control System (DCS). Examples include: main converter roll-in and roll-out sequences; emergency roll-out in the event of power failure; and the charging sequences for reagents (silica and reverts). Other automated features configured for the PSCs were specifically aimed at maximizing converter capacity and reducing the environmental impact from fugitive emissions. These features include: “one button” in-stack charging; automated restriction of blowing rate based on the available acid plant capacity; and fully-automated control logic for the Secondary Ventilation System. In-stack charging allows anode furnace oxide slag to be poured into the PSCs while the converters are still blowing and still sending off-gas to the acid plant. PSC Secondary Ventilation control logic was configured to require minimum input from the operators while ensuring maximum utilization of the available draft from the ventilation system. Highly automated PSC controls contribute to high utilization, averaging 41.2 in-stack hours per day. INTRODUCTION Kansanshi Copper Smelter (KCS) is located near Solwezi at Kansanshi copper mine in Zambia's Northwestern province. The smelter treats concentrate produced at Kansanshi mine as well as Sentinel mine located 150 km west of Solwezi. Four Peirce-Smith Converters (PSCs) are installed at KCS. The nominal capacity of each converter is to treat 350 t of copper matte and produce up to 250 t of blister copper per batch. Normally three PSCs are hot, with two able to operate (produce off-gas), and the third converter on hot standby. The fourth converter is normally under cold repairs. The blister charge size is relatively low for the matte charged due to the unavailability of tank house or other high copper scrap.

Jan 1, 2019

By Garth D. Kirkham

The Problem The Don deposit is a series of two sub-vertical zones that pose a number of challenges with respect to modeling and interpolation. The first challenge is that, based on data and observations, the mineralization and, more importantly, the grade is layered or banded and is stratigraphically controlled. The second challenge is that, due to the undulations of the deposit and extreme changes in dip, using a standard oriented ellipse to guide the estimation process is not appropriate and does not result in an estimate that is representative on a local scale. In order to solve this problem an adapted methodology of relative elevation modelling, has been devised to effectively model and estimate deposits faced with these challenges. Relative Elevation Designed for bedded seam and gridded surface With variable dip and vertical grade variability

May 1, 2013

By J. Ling, J. Xiao, A. R. Laplante

"The most challenging gold gravity separations occur when gangue has a high s.g., gold is fine, or a combination thereof This paper examines such separations, and evaluates the performance of a 3-in Knelson Concentrator at various rotation speeds, a 4-in SuperBowl, and a laboratory Mozley Multi-Gravity Separator (MGS). Synthetic feeds, primary cyclone underflows and flash flotation concentrates are tested.It is concluded that higher rotation speed are mostly useful for very fine (< 3 7 µm) gold recovery, with potential loss of recovery in the 37-212 µm range. This problem is traced to the collapse of the flowing film at higher Gs, which hinders the percolation of particles of intermediate size. Increasing fluidization water flow can alleviate this problem, but the recovery of very fine particle then drops. The MGS does not operate with the same principles, and can achieve good to very good recoveries of very fine to fine gold particles, but the unit has a high cost to capacity ratio, and there is no evidence yet that it can yield smeltable concentrate grade or a concentrate that can be upgraded with a shaking table.IntroductionThis presentation is a sequel to similar CMP presentations of the past eight years; all focused on practical aspects of gold recovery. The first hard rock application of the Knelson Concentrator (Laplante et al, 1990), studies of the grinding and classification behaviour of gold and other malleable metals (Banisi et al, 1991; Laplante et al, 1995), the development of a 3-in Knelson based methodology to study gold gravity circuits (Laplante and Shu, 1992), a comparison of the Knelson and Falcon (Laplante, Laplante et al, 1993), the presentation of a test to determine the amount of gravity recoverable gold, or GRG (Laplante et al, 1996), and the use of synthetic materials, in particular tungsten, to mimic gold (Laplante et al, 1995) were presented. A summary of many practical implication of the McGill research was presented in 1994 (Laplante). Whereas gold gravity recovery was less popular and largely jig-based (in North America) at the beginning of the eighties, many gold plants now incorporate a gravity circuit, and the Knelson Concentrator is the most commonly used unit."

Jan 1, 1998

By R. Tremblay, M. Burrows

"In the summer of 1990, pilot scale testing of a Jameson cell was carried out at Falcon bridge Limited, Kidd Creek Division. The Jameson cell is self aspirating and has fewer mechanical parts, thus it has a lower operating cost compared to conventional cells. This lower operating cost combined with positive pilot scale results, justified the evaluation of two full scale cells in the plant. The plant test had two mandates:1. Gain a thorough understanding of the cells operation and design constraints.2. Develop a flow sheet to replace three divisions of copper cleaner flotation.This paper provides an overview on the installation, commissioning and evaluation of the two Jameson cells acquired. The current status of the test program is also discussed.IntroductionThe Falcon bridge Limited-Kidd Creek Division concentrator is located in Timmins, Ontario. The concentrator, commissioned in 1966 and expanded in 1978, consists of four parallel grinding and flotation divisions utilising small cell (60 ft3) technology. With increasing pressure to maintain low operating unit costs a series of studies were undertaken to assess the impact of new technologies in the flotation circuits. The goals of the project were to improve productivity through increased metallurgical performance and/or reduced operating costs.One of the technologies investigated was the Jameson cell designed by Dr. Graeme Jameson of the University of Newcastle and marketed by MIM Holdings Limited, Australia. This style of column technology was chosen based on its potential for reduced capital and operating cost versus large conventional cells.The success of pilot testwork in 1990 led to the installation of two production sized Jameson cells beginning in October 1991. The units were to be used to confirm pilot results and determine equipment requirements for flotation modernisation. This paper will discuss the results and experiences Kidd Creek has had with the Jameson cell from pilot scale through installation and commissioning of the plant units."

Jan 1, 1993

By M. Redfearn

"DUST COLLECTION EQUIPMENT LIST Primary Dust CollectorModular Mikro-Pulsaire - Model 2G1-96TR Capacity: 20,000 c.f.m.Bags: Total of 288; 8' x 4 1/2"" diameterMaterial: Dacron HCE feltAir to cloth ratio: 7.55/1Bags are changed every 6-9 months. Sent to A & L Services for cleaning at a cost of less than 1/3 the price of a new bag. No new bags have been required since start-up 4 years ago.Blow hack: 35 c.f.m. at 100 p.s.i.Fan: Joy Manufacturing, Model 3300, Type 4S, Backward inclined fan.Impeller: 33"" diameter36"" diameter Inlet33"" diameter Outlet at 3200 f.p.m. velocity1945Motor:,60 H.P., 1780 r.p.m.Capacity: 23,400 c.f.m.Design:,20,000 c.f.m.,•Actual:,17,500-19,000 c.f.m. depending on bag condition.Dust Slurry Agitator: Port-a-flo Model 050-350Major Maintenance:The fan was sped up from 1830 to 1945 r.p.m. and balanced at the higher speed. Several months later the impeller destructed beyond repair. A new shaft and impeller was installed, casings straightened, and bearings reused. Unit is currently running at 1945 r.p.m. No definite cause for the impeller failure has been found, although it might be attributed to improper balancing."

Jan 1, 1976

By J. B. Sabey, V. R. Degner, D. MacGillivray

"Hydrodynamics principals are applied to the design of a natural air ingestion flotation machine for coarse solids applications. Preliminary laboratory test data and subsequent operating plant flotation kinetic performance evaluations are reviewed. INTRODUCTIONThe concentration of potash (sylvite, KCI) from halite and other salts using flotation is well understood and practice a with much success. Reagent schemes and plant circuitry for obtaining the desired product recovery and grade have been established with proven performance. These flotation circuits normally process the medium and finer size particles, less than 30 mesh (600 microns).However, often the potash market trends specifically demand a coarse product size. This emphasis, accompanied by the existing high losses of coarse (+30 mesh) potash in flotation circuits, creates the need for development of a flotation machines specifically designed for coarse potash beneficiation.This technical challenge was undertaken by the WEMCO Research and Development Group. The specific objectives of this flotation machine development program were ( l) to achieve an ""Ultra Coarse Flotation Machine"" capable of adequately suspending and transporting particles in the 4 - 30 mesh (4.75 - 0.6 mm) size range at solids concentrations of 30% or greater and (2) ascertain metallurgical performance of this machine.This paper covers the development (both theoretical and mechanical/ hydrodynamic), the field test program and results, a discussion of applicable flotation kinetics, and conclusions of this ""Coarse Potash Flotation Machine"" development project."

Jan 1, 1986

By Ralph Harris, Javier Jara

"Oxygen is essential in gold cyanidation. It is supplied to the leaching media by sparging oxygen or compressed air into the cyanidation tank. The rate of oxygen transfer is critical to achieve the desired oxygen concentration in solution.The performance of a novel device for gas sparging recently developed at McGill University is compared to the mass transfer characteristics of the two spargers most used in the gold industry: the four lance configuration and the inverted cone. The distribution of gas and the kinetics of oxygen desorption of the three types of spargers was studied in a 140 litre tank. The oxygen dissolution of the new device is three to five times faster than the other spargers. This could increase the leaching rate and/or decrease the cyanide and oxygen consumption.IntroductionIn hydrometallurgical processing, it is often necessary to bubble gas through a pulp to carry out the wanted reactions. The gas can act to agitate the pulp and it can also supply reactant. The rates of the reactions which require a reactant supplied by the gas are often dependent on the rates of mass transfer from the gas to the solution.A common example of this is gold cyanidation which requires oxygen for gold dissolution. Since the gold containing slurry is rich in other oxygen consuming substances (sulphide like pyrrhotite), it is important to provide enough oxygen to allow complete gold dissolution within the residence time of the cyanidation reactors. It has been demonstrated that the rate of gold dissolution is directly proportional to the dissolved oxygen concentration of the pulp which in turn is proportional to the oxygen partial pressure in the bubble 1 2 3. Improving the rate of mass transfer of oxygen increases the dissolved oxygen concentration of the slurry and therefore improves the rate of the reaction.The rate of mass transfer from a bubble to its surrounding solution increases with decreasing bubble size and increasing gas bubble residence time which in turn increases with decreasing bubble size. The generation and dispersion of small bubbles is thus critical to achieving high rates of mass transfer. Different configurations and devices used to inject gas have different effects on the bubble size and the residence time, influencing the overall rate of mass transfer. Therefore better mass transfer can be achieved by adopting a more efficient method.The methods tested in this report are the four lance configuration (wall spargers), the inverted cone and a new device, designated HCD. The wall spargers and the inverted cone are used widely in the industry today."

Jan 1, 1994

By P. Karimi, A. Amini, H. Abdollahi, A. Mirzaeian

Due to detrimental effects of cyanidation process on the environment in gold industry, application of cyanide alternatives such as thiourea are being highly considered in recent projects. In this study, the extraction of gold from poly metal ore, using thiourea was precisely investigated. The representative sample contained 2.5 ppm Au and 160 ppm Ag. The sample was categorized as oxide gold ore, because the main part of its pyrite was changed to iron oxide and hydroxide such as goethite and lepidocrosite. According to mineralogy, the main gangue mineral was quartz. It shows that gold grains are free and interlocked with quartz matrix. Also, the presence of a high amount of silver in free and sulfide forms and also the low content of other sulfide minerals such as base metal sulfides is a remarkable point of these studies. In the gold extraction process with thiourea, there are several parameters that affect the rate of gold dissolution such as concentrations of thiourea as a gold lixiviant and ferric sulfate as an oxidant, pH and size fraction of feed and leaching time. The influence of these parameters on the gold dissolution rate were investigated and optimized. The results showed that the concentration of thiourea is the most influential parameter which could increase the gold recovery to 71% (at 10 g/kg sample). The optimal amount of ferric sulfate was determined as 6 g/kg sample. Also, the optimal levels of size fraction of feed and pH of solution were 88% under 75 micron and 2 as acidity; respectively. Leaching time was 6 hrs. In addition, this sample was leached with cyanide and the optimized values of size fraction of d88=75 microns, pH=10.45, 2500 ppm of sodium cyanide with time leaching of 24 hours were concluded for maximum recovery of gold in solution. Finally, the results of cyanidation tests were compared to the thiourea dissolution after optimizing all affected parameters in cyanidation process.

Jan 1, 2016

By Kim H. Shikaze, John R. Hawley

"Even though acid mine drainage is an old problem, particularly in coal mining areas, its appearance in the form of a major environmental hazard in the hardrock mining districts of Ontario was unexpected. The yellow and amber coloured wastes that characterize an acid mine drainage situation were first noted by the Ontario Water Resources Commission in the uranium mining Elliot Lake district. Problems of a similar nature were also noted in the copper-lead-zinc camp at Manitouwadge, the nickel camp at Sudbury and at numerous isolated base metal properties, both in and out of the Province of Ontario.Without exception, when acid mine drainage type effluents were evident, the specific companies involved worked sulphide or sulphide-associated ores. Without exception, the ore bodies contained quantities of iron sulphide (in various mineralogical forms) and, without exception, the milling circuits serving these operations were such that the iron sulphides were rejected to the tailings areas as waste.Once in a tailings area and after a varying period of time, iron sulphides react to form water soluble salts which, when discharged in the effluents from the tailings area, can affect stream chemistry and stream ecology.It is not unusual that the minerals pyrite, marcasite and pyrrhotite (all iron sulphides) react to form a stable oxide under atmospheric conditions. The law of stability of rocks and minerals, for instance, states that rocks and minerals are stable only in the environment of their formation and, as the environment surrounding them changes, the rocks and minerals tend to change to other species which are relatively more nearly stable under the conditions of the new environment (a). Under atmospheric conditions, iron oxide exhibits a greater stability than iron sulphide."

Jan 1, 1971

By C. -G. Berg, J. T. Tuominen

Since the beginning of the 20th century, energy efficiency in copper smelting has been improved by a factor of 30 as a result of adopting autogenous smelting processes with more and more efficient waste heat recovery systems, increasing use of oxygen enrichment and countless smaller improvements. Nevertheless, energy in the form of electricity and fuels continues to be the biggest contributor to smelter operating cost. Metallurgists around the world are under constant pressure to find ways to save those precious megajoules wherever they can. One area that is often neglected is the removal of water contained in concentrates. Copper flotation concentrates, as delivered to a smelter, typically contain some 8–12% moisture. This moisture must be removed and the only way to do this is by evaporation, which consumes a lot of energy. Several studies have shown that the most efficient way of removing this moisture is to use a steam- heated indirect dryer—the least efficient way being the feeding of wet concentrate directly into a smelting furnace. Even in the most efficient smelters that employ modern steam drying, the energy needed to remove the moisture from the concentrates constitutes over 15% of the total energy consumption of a smelter – acid plant complex. The major steps that have been taken in the past to lower the energy consumption of moisture removal are the use of drying in general and the adoption of steam drying. The energy efficiency of concentrate drying has been improved by 25% over the last 20 years. The latest giant leap in this area, the Kumera Heat Recovery Dryer, is the result of Kumera’s long-term development in improving the efficiency of concentrate drying. The Heat Recovery Dryer is a novel application of steam drying that uses the waste heat from dryer exhaust gas for drying concentrates, thus, cutting the energy consumption by a further 30%. This paper discusses energy consumption in copper smelting in general and analyzes the energy consumption of concentrate drying in detail. Different dryer types are compared especially from the perspectives of energy consumption and environmental emissions. Applicability of different energy sources for drying is evaluated and discussed. The principle of the Kumera Heat Recovery Dryer is presented along with results from extensive continuous pilot plant tests. Potential applications and the feasibility of Kumera Heat Recovery Drying technology are discussed.

Jan 1, 2019

By A. Mukherjee, C. Kelly, B. Markovich, G. E. Agar, G. E. Robertson, F. Khan, B. Shea

"Laboratory batch flotation tests were performed on a sample of bulk slow cooled converter matte (Cu:Ni = I) to optimize the time in each of the separation stages. Numerical simulations were carried out to estimate a continuous circuit material balance for the conditions selected. An experimental simulation of a continuous circuit (locked cycle test) readily came to steady state and there was reasonable agreement between the numerical and experimental results. The use of nitrogen as the flotation gas and the addition of collector to the grinding mill minimized the oxidation of chalcocite, which has previously resulted in increasing circulating loads, and led to a stable circuit with very low weights in the recycle streams.IntroductionINCO Limited developed a flotation process for separating the components of copper-nickel matte [Sproule, Harcourt and Rose 1945; Sproule, Harcourt and Renzoni 1960] in 1945 to replace the Orford process and continues to use the flotation process in almost the original mode to this day. Initially the matte processed was the result of smelting a nickel concentrate that analyzed about 12% nickel and 2% copper and all of the recycle material that was generated in the smelting process. The matte generated analyzed about 25% copper and 53% nickel, roughly 2: 1 nickel to copper. This matte was blown in Pierce-Smith converters to be sulfur deficient in order to produce a metallic phase that collected the precious metals contained in the smelter feed [Wilson and Stratton-Crawley 1991]. This process continued with little change until 1994 when the separation of chalcopyrite from pentlandite in the Copper Cliff mill was discontinued after 63 years of operation. As a consequence of the elimination of the primary separation of chalcopyrite from pentlandite in the milling process a bulk concentrate was produced for smelting. Bulk concentrate smelting came into practice in 1994 as part of the initiative to meet government legislation that restricted the emission of sulfur dioxide to the atmosphere. Previously, the copper concentrate was smelted in an INCO oxygen flash furnace which resulted in the capture of a high strength sulfur dioxide stream from which liquid sulfur dioxide was produced. Nickel concentrate was smelted in reverberatory furnaces that produced low strength sulfur dioxide gases which were difficult to treat. The action taken by INCO to meet the environmental constraints was to smelt a bulk concentrate in oxygen flash furnaces which would produce gases suitable for the production of liquid sulfur dioxide and another stream that was high enough in sulfur dioxide to be readily converted to sulfuric acid [Bell et al 1990]. The end result of this decision was that all the copper-nickel separation would be done in the matte separation plant."

Jan 1, 1996

By P. M. STILES

Acquired in 1957, the limestone quarry of Ideal Basic Industries on Texada Island north of Vancouver has become a major supplier of high purity limestone and construction stone. Crushed limestone products are shipped by barge to various destinations along the B.C. and northern U.S. coasts.

Jan 1, 1984

By S. Masambi, S. M. Bradshaw, C. Dorfling

The presence of iron in base metal leach solutions complicates nickel and copper recovery processes. Leach solutions are commonly purified by precipitating iron as hematite, jarosite or goethite. Fast oxidation of ferrous to ferric is an important step to remove iron by precipitation at low pH values. This study reviews the ferrous oxidation mechanism and reaction rates in chloride media using molecular oxygen. This work contributes to the development of a better understanding of iron precipitation processing steps, which continues to be an important matter in hydrometallurgy. The specific objectives were to investigate (1) the effect of key process variables (acid concentration, copper concentration, and temperature) on the rate and extent of ferrous oxidation by dissolved oxygen and (2) the oxidation mechanism by varying chloride and hydrogen ion concentrations separately. Batch oxidation tests were performed using synthetic chloride solutions containing 45 g/L Fe and 3 g/L Ni at 50 and 80°C, dissolved Cu concentrations (0.3 to 3 g/L) and HCl concentrations (2 to 4.6 M). Regression analysis indicated that the significance of the effect of different variables on the extent of oxidation achieved after five hours decreased in the following order: HCl concentration > Cu concentration > temperature. Increasing the HCl concentration from 4 M to 4.6 M caused the degree of oxidation achieved after two hours to increase from 58.3% to 98.1%. The catalytic effect of dissolved copper on ferrous oxidation was more significant at low acid concentrations. Experimental data suggest that rapid oxidation of ferrous is possible at high HCl concentrations for two main reasons. First, the formation of ferrous chloride complexes allows ferrous oxidation to proceed without the formation of superoxide as an intermediate product (the rate limiting step in dilute acid solutions). Second, the formation of ferric chloride complexes stabilises the oxidation product.

Jan 1, 2016

By J. Donovan

Underhand cut and fill mining is being planned for selected areas at the David Bell Mine where high ground stresses have resulted in particularly challenging mining conditions. In order to test the placement and performance of a variety of underhand cut and fill paste mats three different designs have been installed and extensively instrumented along the A-Zone 9C ore sill. The sill is approximately 87m long, 4.0m high, and 5.0m average width. The sill was broken into 3 sections each of which will contain one of the sill mat designs. The underhand sill will be mined 3.75mW x 4.0mH, allowing for some overbreak while minimizing the risk of the back span exceeding the design limit of 4.0mW. Each test sill mat includes pre and post-support and includes 0.5m (18?) broken muck mat placed along the sill floor. Upon developing beneath these mats quantitative and qualitative measures will be used to determine which design is the most suitable for application in underhand mining at David Bell.

May 1, 2007

By B. R. Neufeld, M. G. Heximer, S. J. Roach, W. G. Van Luven, D. G. Reynolds

"Teck Metals Ltd. operates an integrated zinc-lead complex at Trail, British Columbia, Canada. In the zinc plant, zinc concentrates are roasted to produce calcine, which is then processed in the leaching circuit for zinc recovery. Calcine is first dissolved at pH <2.0. This is followed by a precipitation stage using additional calcine that removes iron as a ferric hydroxide precipitate which is recycled to the lead operations. It is during this pH adjustment that impurities, notably arsenic, antimony and germanium, are co-precipitated with ferric hydroxide which is the primary removal phase for these elements. Recent processing conditions have resulted in lower iron inputs to the leaching circuit and this has resulted in insufficient co-precipitation of key impurities. This paper summarizes the influence of iron concentrations on impurity removal using laboratory and plant scale surveys, and emphasizes key specification limits for the primary iron removal circuit.INTRODUCTIONThe Sulphide Leach Plant (SLP) at Teck Metals Ltd., Trail Operations produces a zinc sulphate electrolyte from three different zinc bearing streams:1. Calcine, primarily containing zinc oxide, from the roasters2. Zinc electrolyte from the Oxide Leach Plant (OLP) is produced by leaching zinc oxide fume from the pyrometallurgical lead smelting operations3. Zinc slurry from the Zinc Pressure Leach (ZPL) autoclaveThe feed sources are blended and introduced to the iron removal process to precipitate iron in the form of a hydrolyzed iron precipitate which is recycled to the lead smelting operation. Along with iron, impurities, namely arsenic, antimony, germanium and silica, also co-precipitate; this process acts as the primary removal step for these elements. Following iron removal, the electrolyte is further processed in the zinc dust purification circuits. Cadmium and copper are removed in the cold stage circuit via zinc dust cementation, followed by cobalt cementation in the hot stage circuit."

Jan 1, 2016

By P. D. Duever, P. J. Readyhough, P. L. Silveston

"The Plastojrost apparatus, developed by Ritter and Juranek in the 1960s has been modified primarily, to automatic temperature control. A limited investigation of carboniferous and cretaceous coal samples showed that both coals undergo the same changes when coked in the Plastofrost unit. Dilatometer plastic range and total dilatation correlated well with the Piastofrost plastic range for the carboniferous Appalachian coal, but not for the cretaceous western Canadian coal. Work to elucidate the characteristics of some western Canadian coals using the Plastofrost is in progress. IntroductionCoke quality is critically affected by the properties of the plastic phase in the coal-to-coke transformation. Since these properties exist in a coke oven and are not measurable, they are approximated by standard laboratory tests, such as free swelling index, plastometry, and dilatometry. Of particular interest to us is the anomalous behaviour of western Canadian coals of cretaceous origin. By many predictive methods, these coals should not produce coherent cokes. When coked in pilot scale test ovens, in fact, they produce cokes of good stability. They typically make up 20% of Japanese coking coal blends and are also used in South Korea, Mexico, Brazil, and several European countries. This anomalous behaviour has been studied in our laboratory<8.9.1O). Plastic properties or rheological tests examine just the external ramifications of the coal-to-coke transformation. To understand the process, however, microscopic examination of the coal and coke seems to be needed. The Plastofrost unit has been shown to be a convenient means to this end."

Jan 1, 1987

By M. Pelser, H. Gous, J. H. W. M. Brits, K. Voogt

The effect of impurities; iron and selenium, on current efficiency and deposit quality in copper electrowinning was investigated using an at-line cell at the Rustenburg Base Metals Refiners’ Copper Tankhouse. The tankhouse is a direct leach, starter sheet electrowinning operation with a unique electrolyte composition. The effect of iron on the current efficiency in the at-line cell was found to be less severe than reported in literature. However, the distinction between the chemical impact of iron on current efficiency and the physical impact of iron on a starter sheet operation needs to be realised. The effect of selenium on current efficiency and deposit quality is also reported. INTRODUCTION Anglo American Platinum’s Rustenburg Base Metals Refiners’ (RBMR) Copper Tankhouse serves a unique function in the refinery flowsheet in that its primary functions are the removal of copper and production of acid for the leach circuit. This function needs to be performed to specification in order to keep the leach circuit operating optimally. Current efficiency is seen as an indicator of tankhouse performance and efficient energy use. The tankhouse plates starter sheets onto titanium blanks which, after stripping and preparation, return into the production cells for further plating. Approximately 78% of tankhouses surveyed (Robinson et al., 2013) use permanent cathode technology whereas older, smaller operations still use starter sheet technology. Drawbacks of the starter sheet technology are that the starters are thin and bend easily leading to increased shorts, they need to be removed after two days of plating to be manually straightened and are then returned to the cells. In addition, the electrolyte levels in the cells need to be adjusted during the plating cycle to improve loop adhesion to the cathode and reduce loop corrosion. The starter sheet technology is therefore very labour intensive and the final quality of the product is dependent on the quality of the starter sheets produced. In comparison, permanent cathode or direct plating technology lends itself to automation, allows for higher current densities and better cathode quality due to more robust geometric stability in the cell. The present barrier to implementation of permanent cathode plating at the RBMR Copper Tankhouse is the capital cost required to convert the tankhouse, i.e. blanks and stripping machine. At present, alternatives to titanium blanks, such as LDX 2101, are being investigated for the unique electrolyte composition of this tankhouse.

Jan 1, 2019

By Lawrence Devon Smith

"The concept of risk is well established in the mining industry. It is acknowledged when estimates of reserves are expressed as proven or probable, when mining and metallurgical recoveries are applied to the ore, and when contingencies are added to costs. However, it is difficult to provide a quantitative assessment of risk. The significant sources of risk in a mineral project and how these are addressed in project evaluations are discussed.The discount rate is examined as a fundamental means of reflecting risk in discounted cash flow evaluations. Current industry practice is discussed, and a methodology for the analysis of risk levels is proposed that assesses the constituent components of the discount rate: real interest, mineral project risks, and country risk.IntroductionThis paper examines the components of the discounted cash flow discount rate and proposes a method of estimating project-specific discount rates. The motivation for this work comes from experience of project evaluations in which owners and purchasers had agreed on virtually every aspect of the evaluation except how much the project was worth. Agreement was reached on reserves, grade, recovery, capital costs, operating costs, taxes, and, by combining these components, even on the final cash flow values. The only difference in opinion concerned the discount rate to be used in the calculation of the net present value. Depending on the life of the project, such differences of opinion can cause a variation of more than 50% in the value placed on a project! This is illustrated in Figure I."

Jan 1, 1995

By J. C. Taylor, J. A. Davis

"Existing environmental, health and safety regulations have forced pyrometallurgical facilities worldwide to drastically change some of their traditional methods of operation. In extreme cases, a number of facilities have been closed due to their inability to conform to these existing regulations. Tougher future legislation and government policy will lead to more stringent regulations and enforcement policies. The efficient control of smelter secondary emissions will, therefore, become critical to successful smelter operation.Much attention has been paid to the smelter primary process off-gas systems, including the development of new process designs to reduce the number of uncontrolled transfer operations. However, even the most modem smelter design requires control of its secondary contaminant sources, which mayor may not be directly related to the process operation.This paper discusses some of the potential secondary emissions encountered in many of the present process technologies. A few of the emission sources presented are only periodic in nature, such as those caused by maintenance procedures, but which have the capability of forcing smelter shut-down because of their direct impact on worker health and safety. The paper deals with a variety of topics relating to the efficient control, maintenance and monitoring of secondary emissions in copper, nickel, aluminum, magnesium and lead smelters, as well as in iron foundries and steel making facilities. Examples of the typical problems encountered in each of the above process operations are presented, including fume control at furnace tapping and slagging locations, metal transfer stations, tapping launders, and ladle transfer operations. The parameters for the most efficient design of fume capture systems, problems encountered with the handling and clean-up of certain collected emissions, and examples of successful solutions in particular operations are discussed."

Jan 1, 1994

By K. Hashiba, Y. Z. Liang, K. Fukui, M. Koizumi

"Efficiency and penetration rate during percussive rock drilling significantly affect the speed of tunneling and excavation. Many researchers and engineers have continuously engaged in the advancement of rock drilling performance. Development and modification of the rock drills based on test data are timeconsuming and costly. Hence it is essential to make effective use of numerical simulation. However, most of the numerical models proposed in previous studies were too simple in comparison with the actual percussive rock drilling in which many factors are involved. The objective of our research project is to construct a new realistic numerical model of continuous percussive drilling, which is based on the previous Okubo-Nishimatsu’s model. In this study, the parts of the previous model involved in stress wave propagation and attenuation were modified. The two models were prepared for HD712 hydraulic rock drill mounted on crawler drills and HD210 hydraulic rock drill mounted on drilling jumbos. Stress wave propagation was computed with the discretized one-dimensional wave equation. The calculated results were compared to the stress waves measured on a rod during percussive drilling. In both models for HD712 and HD210, the measured stress waves can be reproduced with the one-dimensional model of the same acoustic impedance as the actual piston, shank rod and rod combined with the CI+spring model as a rod joint. The calculated results indicate that the spring constant in the model of a rod joint needs to be set appropriately depending on drilling conditions. INTRODUCTIONEfficiency and penetration rate during percussive rock drilling significantly affect the speed of tunneling and excavation. Many researchers and engineers have continuously engaged in the advancement of rock drilling performance. Components related to the percussive drilling are a drill body, a piston, a shank rod, rods, rod joints, a bit and rock. The piston in the drill body is reciprocated by hydraulic pressure and collides with the shank rod a few thousand times per minute. Stress waves generated by the collision propagate and attenuate in the rods and rod joints, and then reach the bit. The bit penetrates and crushes rock. During percussive drilling, the components mutually affect each other and self-induced oscillation occurs. The number of blows and the rate of penetration are determined by the practical drilling conditions such as hydraulic pressure, thrust force, number of rods and mechanical properties of rock. Development and modification of the rock drills based on test data are time-consuming and costly. Hence it is essential to make effective use of numerical simulation. However, most of the numerical models proposed in previous studies were too simple in comparison with the actual percussive rock drilling in which many factors are involved."

Jan 1, 2015