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By R. Haecker, S. Bajic, R. Thorpe, A. Palko, M. Kalstrom

"Historically, mineral processors have used inferential data like plant instrumentation, assays, and grind sizes to monitor the performance of their plants. The ability to monitor the minerals themselves on a daily basis has eluded all but the very largest and richest of operations. The advent of a new generation of compact, robust, and user-friendly SEM-based automated mineralogy systems looks set to accomplish this, aiming to bring daily mineral monitoring of plant performance within the reach of all mills from mid-size upwards. This paper describes the use of this machine at Greens Creek Mine, a 2.5 ktpd Ag-Au-Pb-Zn operation with a complex flotation flowsheet. During the trial, the machine provided turnaround from sample collection to reporting averaging 24-36 h for key daily samples. All equipment was located within the compact mill, and proved highly reliable despite the challenging environment. Turnaround time was fast enough for metallurgists to understand, and, where possible, respond to ore and process changes while they were still happening. The plethora of data generated was filtered to a simple report whereby mill management could quickly absorb key daily mineral-based performance indicators. This paper will describe how the data was used to understand the key drivers behind metallurgical efficiency at Greens Creek, and how metallurgists began using the information to enhance daily mill performance."

Jan 1, 2014

By Sune Eriksson, Stig Peterssonsune, Christer Fridfeldt

"Boliden Metal Corporation has developed a technique to treat complex copper concentrates containing large amounts of lead, zinc, arsenic, antimony and bismuth which cannot be treated in conventional smelters due to the impurities.Tests were carried out to treat such complex copper concentrates in the TBRC. They were performed in a pilot-plant TBRC with charges of 5 tons. These tests resulted in a new method to treat complex copper raw materials and produce a blister copper low in impurities. The background and data from the test work are presented as well as a flowsheet for a full-scale process.IntroductionThe Ronnskar Works is situated on the coast of the Gulf of Bothnia, 40 miles east of the Boliden Mine. Smelting began at Ronnskar in 1930 to process the complex ore extracted from the Boliden Mine. The Ronnskar Works has subsequently developed into a relatively integrated smelting plant.Copper and lead sulphide concentrates from Boliden's own mines as well as purchased copper raw materials are processed. The main products are refined copper and lead, but full utilization of the raw material also yields other metals and some metallurgical and chemical by-products. The products include gold, silver, zinc clinker, white arsenic, arsenic salts and arsenic metal , selenium and selenium compounds, nickel sulphates, sulphuric acid and liquid sulphur dioxide."

Jan 1, 1981

?Why TSM in Canada ?How it was developed ?Architecture and Goverance ?COI panel ?Performance indicators and how they work ?External Verification ?Reporting ?The national organization of the Canadian mining industry ?Mining, smelting and refining of base and precious metals ?Mining of iron ore, diamonds, uranium, oil sands, coal and industrial minerals ?Many associate members ?www.mining.ca

May 1, 2011

By Jean Cayouette, Paul Blatter, Paul Cousin

"The Laronde mill was commissioned in 1988 and has been treating the Laronde Penna ore since 2000. The Laronde plant, situated in the Abitibi region of Quebec, has experienced mill expansions to accommodate 7500 mtpd and to recover gold and silver as well as important amounts of base metals (Cu, Zn, Pb). Ten sequential metallurgical and environmental processes are responsible for metals recovery and process viability. Over 200,000 ounces of gold, 4,5 million ounces of silver, 7000 tonnes of copper, and 78,000 tonnes of zinc are produced annually.In 2008, the Goldex plant began trucking high grade (35-45 g/t Au) pyrite sulphide concentrate (TCGO) for gold extraction at a newly constructed treatment plant, increasing gold output at Laronde by 40,000 ounces. In 2009, the high gold grade Lapa ore processing facilities were annexed to the Laronde plant increasing output by 120,000 ounces of gold. TCGO and Lapa productions are forecasted until 2018 and 2015 respectively, with known reserves of each mine.Today, up to 140 permanent employees run and optimise the Laronde plant operation. As a result, Laronde has intricacies intrinsic to a large and varied operation. The paper will outline the metallurgical highlights, challenges, and future outlook related to the treatment of each ore type (Laronde, TCGO, and Lapa).INTRODUCTIONFollowing the 2000 to 2003 expansions, the Laronde plant evolved into a mature milling complex with developed processing, man power, and water treatment capabilities able to accommodate the complex Laronde ore. The ore requires a series of grinding, copper/lead flotation and separation, zinc flotation, zinc tails precious metals leaching, followed by a counter-current decantation circuit and Merrill Crowe precipitation. A paste backfill and SO2-air cyanide destruction operate intermittently. The tailings area has a dedicated cyanide destruction and metals precipitation plant prior to water recycling to the mill. A biological water treatment plant was put into service in 2004 to address the build-up of thiocyanate in the tailings ponds at Laronde. This was the result of the high sulphide content of the Laronde ore and 90% recirculation of the process water. The plant uses bacteria to oxidize and destroy thiocyanate, and removes phosphate from the water before it is released to the environment."

Jan 1, 2011

By John Lindsay

Outline ??Recovery plant requirements and concepts ??Common recovery plant unit processes ??Recovery plant design considerations ??Peripheral issues Process Design ??Liberation ??Separation ??Recovery Processing ??Density 3.5 g/cm3 ??Luminescent under x-ray ??Hydrophobic (non wettable) surface ??Non-magnetic ??Inert to strong acids / bases

Apr 1, 2005

By R. H. Walton

"In recent years the recovery of copper from low grade oxide ores has received widespread attention. The cost of treating such ores using conventional technology, such as milling and flotation, is frequently prohibitively high and copper recovery is generally low. The introduction of heap leaching, solvent extraction (SX) and electrowinning (EW) has resulted in high recoveries, good copper cathode quality and low production costs. This technology has now been successfully applied on a worldwide basis.This paper reviews the various process and design options and gives approximate capital and operating costs for a 1.0% copper ore.INTRODUCTIONAs a result of the low copper prices that prevailed for much of the last two decades, aggressive attempts were made by the industry to cut operating costs. Under such circumstances heap leaching, solvent extraction and electrowinning, to recover low cost copper from oxide ores, has received particular attention and ultimately widespread acceptance.Low grade copper ore (typically containing less than 1.0% Cu), which could not be economically treated by other methods, is placed on heaps and leached with dilute solutions of sulphuric acid. The copper rich, pregnant solution, drains from the heap into a collecting pond and is pumped to a solvent extraction plant where it is purified and upgraded. The resulting solution, termed the pregnant electrolyte, passes to the electrowinning section where the copper is recovered as high grade cathodes and sulphuric acid is regenerated, (Figure 1).This paper specifically describes the processing of copper oxides. Other copper minerals such as chalcocite and chrysocolla are also leached by dilute solutions of sulphuric acid. In practice the oxide minerals are frequently associated to varying degrees with less easily leached sulphides, such as chalcopyrite. In some cases, once the oxidized capping of the ore body is removed, the oxides are totally replaced by more refractory sulphides. Under these circumstances the leaching recovery is significantly reduced. However the principles described here for solvent extraction and electrowinning are also applicable to solutions received from heaps containing sulphide minerals, which are leached more slowly and involve bacterially induced reactions."

Jan 1, 1990

By S. Hunt, D. de Vries

Kansanshi Copper Smelter began operation in March 2015. The smelter features a single Isasmelt furnace for smelting, one 6-in-line electric furnace for matte-settling and slag-cleaning duties, four Peirce- Smith converters and two rotary anode furnaces. A 1,450 t/day oxygen plant and a 4,085 t/day sulphuric acid plant are included in the smelter operations. This paper gives an overview of the operational performance to date with summary data for each smelter section. Challenges such as an unforeseen refractory failure in the Isasmelt furnace in 2016 are discussed. Design improvements around the electric furnace are mentioned. Plans for potential future improvements are put into context. A consistent focus on operations, maintenance and continuous improvement has seen the smelter easily achieve its nameplate capacity of 1.2 million tonnes per annum of copper concentrate throughput. The smelter now operates consistently at a concentrate throughput rate of 1.38 million tonnes per annum of copper concentrate. INTRODUCTION The smelter was commissioned in March 2015. Design and commissioning of the smelter have been discussed by de Vries, Hunt, Hanschar and Louie (2016). The smelter rapidly ramped up to achieve nameplate production (1.2 million tonnes per year concentrate) and in 2018 operated consistently at the rate of 1.38 million tonnes per year concentrate. This paper presents major operational data for each section of the plant and describes significant events and improvements during the first four years of operation. The smelter flowsheet is shown in Figure 1.

Jan 1, 2019

"A simple approach to optimizing the performance of an operating carbon in pulp (CIP) or carbon in leach (CIL) plant is presented. It is possible to reproduce the performance of an operating CIP or CIL plant in the laboratory, in small scale batch tests in which the rates of gold leaching and carbon adsorption are determined, coupled with mathematical models that translate the batch kinetic data into predicted steady state profiles of gold in solution and on the carbon across multi-stage adsorption circuits.Once the performance of the commercial plant has been mimicked by the model, the plant operating parameters can be varied systematically in the model, and the likely response of the plant to the stepwise change can be determined. In this way it is possible to inexpensively optimize the design of both greenfields CIP or CIL circuits and plant expansions, or optimize the operating parameters of an existing plant, with no disruption to plant operations and no economic risk to the mining company.This approach has been adopted by IAMGOLD Corporation, to the immediate benefit of two of their existing gold operations. The methodology adopted and the results achieved are presented.INTRODUCTIONThe carbon in pulp (CIP) and carbon in leach (CIL) processes became firmly established in the gold mining industry in the 1980s, initially in South Africa and Australia, from where they spread rapidly to all the gold producing regions of the world. The percentage of annual global gold production by activated carbon-based processes grew from zero in the 1970s to almost 70% by the turn of the century, which represented a phenomenal rate of acceptance of a new technology by a traditionally conservative industry. The main reason for this rapid acceptance of the new technology was the fact that the first few large plants built in South Africa convincingly demonstrated better gold recoveries than the traditional filtration/Merrill Crowe process, with lower capital and operating costs. And as the plants developed an operating track record over their first few years of life, they proved to be remarkably robust mechanically, and highly tolerant of plant upsets, changes in feed composition and solution phase contaminants that had caused problems in Merrill Crowe plants."

Jan 1, 2011

By Hu Harries

WE RECENTLY completed a study for a client, who is an applicant to export gas from the Province of Alberta, on the effects of gas export on Alberta's economy. This study was part of the client's application ?before the Conservation Board, and, no doubt, some of you are familiar with it. Submissions by other applicants in earlier hearings before the Board and before the Royal Commission on Energy, as well as claims made by parties opposing export, have also brought forward a number of other opmwns regarding possible economic effects. Ali of these studies have been on a piecemeal basis. They have dealt with particular aspects of particular applications, and only indirectly and by inference with the overall question of gas ex-port. We have prepared this paper to review the overall picture and ex-amine, as far as time permits, all sides of the question. While the paper is entitled Economic Effects of Gas Export on Alberta, we felt that the area considered should in-elude northeastern British Columbia, not only because Westcoast, the first major gas export project, is essentially a northeastern B. C. project, but because this area has historically been, from the point of view of economic, if not of political, geography, part of Alberta. In this paper, when we talk of export, we mean export from the Province or, in this case, from Alberta plus northeastern B. C. Sales to other parts of Canada have exactly the same effect on the economy of this area as l1ave sales to the United States, provided, of course, they are made on equal terms.

Jan 1, 1958

"Geoscan is used for multi-elemental analysis of conveyed bulk materials in the minerals industry with proven performance and demonstrated paybacks of weeks or months in iron ore, manganese, copper, zinc-lead and phosphate industries. Geoscan-M measures continuously through the full bed depth and is unaffected by belt speed, particle size, layering, mineralogy and dust levels. Increments of typically two minutes are accurately measured and provide the plant operators with tonnage weighted results of contained elements. Measurement of the composition in real time enables prompt feedback to the mining operators and also feed forward to the process operators. Crushed ore increments can be diverted based on composition, such as diverting waste for disposal or marginal material to a stockpile for separate processing. Bulk sorting allows material to be placed into relevant stockpiles for subsequent blending to achieve consistency for process feed. Parameters used for decisions may be primary metal content, deleterious component content, other process parameter, or a combination of these. Process parameter may be the elemental ratio between critical components, e.g. Cu:S ratio, basicity, or the content of an “indicator” element (such as Si indicating material hardness). Benefits will be discussed based on case studies in copper and iron ore.IntroductionAs operators of process plants strive to remain competitive as capital and operating costs increase and commodity prices fluctuate, more focus has been directed at incremental performance improvements at incremental cost rather than major expansions to increase output. The philosophy of working smarter allows plants to adopt technologies proven in other industry sectors to minimise risk. This paper discusses one such technology that the minerals sector is adopting and has found to be safe, reliable and accurate with paybacks of weeks in most applications. New applications continue to be found and major benefits realised at various stages in the processing and handling of ores and products."

Jan 1, 2016

By Adam Moore, Peter Radziszewski, Matt Gallimore

"Stirred mill technology dates back to 1928 where the idea to use “an agitator and spherical grinding media” was presented. In the context of the mining industry, interest in stirred milling technologies has grown over the last 20 years where the typical observation reported inDespite these developments, the tumbling mills are continuing to be used in fine and ultrafine grinding applications. The main reason for this, as communicated at the Canadian Mineral Processors roundtable discussion in 2013, is that the comfort level in understanding stirred milling technologies is not at the level of that of tumbling mills.At CMP 2014, one contribution was made to explore how mill design affects power and energy in a stirred mill. While the 2015 CMP paper explored the application of one scale-up methodology to the Gibraltar case. For the CMP 2016, the aim of this paper is to contribute to increasing the general comfort level with stirred milling technologies by developing a better understanding of the different scale-up methodologies used. This will be accomplished by first completing an overview of the scale-up models found in the literature followed by a review of the different testing and scale-up procedures. This review will be supported by a number of case studies found in the literature. The paper will close with a discussion on some of the challenges that these models and methodologies have and how future research might contribute to overcoming them.IntroductionStirred mill technology dates back to 1928 where the idea to use “an agitator and spherical grinding media” was presented (Stehr, 1988). The year 1948 saw the successful application of Du Pont’s “sand mill” for pigment grinding. Subsequent years have led to the development of different types of stirred mills spanning a number of industries that require fine and ultrafine grinding."

Jan 1, 2016

By E. Amini

"While methods exist to allow the simulation and optimization of existing flotation operations, there is a deficit in the ability to predict the performance of a new ore type pre-processing. A laboratory scale batch flotation testing methodology has been developed to assess the flotation potential of ores (known as the JK Floatability Index or JKFI). Previous work conducted by Alexander and Collins (2008) showed that the test could be used to predict full scale results in base metal ores. In this work, JKFI tests were applied to three different types of ore and one blend from a greenfields gold deposit to show the test applicability to flotation of precious metals. The results from the JKFI test on one ore type were then used to predict the response of all ore types (including the blend). To validate the predicted kinetic parameters, a comprehensive survey was performed on a pilot plant where each of ore types was processed in the circuit. An acceptable agreement was found between the experimental recoveries and predicted recoveries which indicate the potential of this method for use in predicting the flotation response of an ore.INTRODUCTIONOver 100 years, flotation has been considered an efficient method to concentrate different types of sulphide and oxide minerals and coal. Thus, numerous research studies have been undertaken to find an appropriate model to predict the behaviour of particles in flotation circuits (Barnwal et. al., 2006; Bascur and Herbst, 1983; Cecile and Villeneuve, 1991; Deng et. al., 1996; Zheng et. al., 2006). Most methods are based on the ore flotation response in an industrial plant which is expensive and time consuming to characterise. It is difficult to reliably predict flotation response for new types of ores in industrial scale plants.The floatability component modelling (FCM) approach, developed within the Australian Minerals Industry Research Association (AMIRA) P9 project, has been used to derive models of operating flotation plants for the purpose of simulation and optimisation (Alexander and Morrison, 1998; Alexander et. al., 2000; Varadi, 2006). This approach has been applied at a large number of operating sites treating a wide variety of ores. The approach requires metallurgical surveys to be undertaken in conjunction with batch flotation tests and cell characterisation. This allows ore and machine parameters to be derived, which are then used in the P9 Flotation Model. The development of a FCM generally requires a large commitment of resources and time. While there has been significant support of this approach in industry, the FCM approach is limited to ores already being processed (Runge, 2004; Wightman, 2006)."

Jan 1, 2009

By F. E. du Plessis

"The distinction between data suitable for process control and data subject to aliasing only becomes quantifiable with the availability of rapid real-time measurements. Expecting the dynamics of flotation to vary for different ore types, process flows and equipment types; a study was conducted to compare the dynamics of different flotation processes. The results shed light on the critical sampling intervals required for process control which in turn allows process stabilization and optimization. The technology used for the in-line measurements is based on diffuse reflective spectroscopy and has been introduced into the mineral processing industry over the past ten years. It has been widely implemented in the sulphide flotation industry in South Africa for the real-time measurement of grade.BACKGROUND AND OBJECTIVE A study was conducted by Du Plessis and Keet (2010) to quantify the minimum sample frequency required for grade and/or impurity control. The study was based on chromite data collected at 15 second intervals from the concentrate line of a flash float cell in the Platinum industry. The data captured the natural process variation of the entrained chromite through water recovery due to changes in feed parameters and flotation performance. It was found that the minimum rate of sampling required from the studied flash float concentrate was 2 minutes. This sampling interval time accounted for 90% of the variance in the process data. It was further found that a sampling interval time of longer than 4 minutes produced data with no significant correlation to reality. The dynamics of the flotation process has been misjudged in the past but can be quantified with the availability of real-time grade measurement based on diffuse reflective spectroscopy. The objective of the study was to compare the dynamics of different sulphide flotation processes through the comparison of maximum sampling interval times required to combat aliasing. All data were collected from single stream analyzers that are currently installed at mineral concentrators. The technology used is based on diffuse reflective spectroscopy."

Jan 1, 2012

By L. Flett, S. Woods, C. Deredin

"Following a significant reduction in world Imperial Smelting Furnace (ISF) capacity from 2003 to 2008, a major optimization initiative was successfully carried out on the lead upgrading and bulk (zinclead) flotation circuits in the Brunswick Concentrator.This paper describes the process improvements that have taken place in order to reduce the quantity of bulk concentrate:•,Increase in lead upgrading rougher flotation circuit capacity•,Installation of a closed-circuit bulk regrind utilizing MillPebs and 150 mm Cavex hydrocyclones•,Installation of two additional cleaner flotation stages on the bulk circuit•,Optimization of the manipulated variables in the lead upgrading and bulk circuitsINTRODUCTIONThe Brunswick Mine (Xstrata Zinc Canada) is located near Bathurst, New Brunswick and is one of the world’s largest lead/zinc mines. Milling of ore commenced in 1964 from the No. 12 orebody, with an expansion in 1966 to accommodate ore from the No. 6 open pit mine. The planned closure of the mine will take place at the end of the first quarter in 2013. The ore is a complex massive sulphide, consisting of approximately 20% valuable minerals, 60% pyrite and 20% silicates.A simplified flowsheet for the concentrator is given in Figure 1. The overall flowsheet involves SAG milling, secondary ball milling, selective flotation, and dewatering by thickening, filtering and drying. The complexity of the flowsheet evolves from the requirements for fine grinding and the production of four concentrates: zinc (sphalerite), lead (galena), copper (chalcopyrite) and bulk (sphalerite plus galena).Primary hydrocyclone overflow feeds the copper-lead aerator-roughers for flotation at pH 9.0 with soda ash. Plant xanthate (80-20 sodium isopropyl and potassium amyl) and AP407 are used as collectors in the copper-lead circuit. The rougher concentrate is directed to the copper-lead regrind circuit and the tailings are fed to the secondary grinding circuit. The secondary hydrocyclone overflow from the secondary grinding circuit feeds the copper-lead rougher-scavenger banks. Rougher concentrates are reground to a P80 of 28 µm prior to three stages of copper-lead cleaning. The first stage of cleaning is done in an open circuit with the tail reporting to the cleaner scavenger. The cleaner scavenger concentrate reports to the regrind circuit and the cleaner scavenger tail combines with the rougher/scavenger tail to feed the zinc circuit. The second and third stages of cleaning are performed in closed circuit with the cleaner tail being recycled to the copper-lead regrind circuit. The average final concentrate grades 23% Pb, 3% Cu and 704 g/t Ag, and is fed to the copper separation circuit."

Jan 1, 2013

By Olli Aulio, Outokumpu Oy

"GENERALThe Outokumpu Company is today an important European producer of metals from its own mines: copper nickel cobalt, ferro chrome and zinc being the main products. In 1972 about 4 million tons of ore was hoisted from eleven mines.Outokumpu Oy has gained experiences on the various aspects of on line computer control since 1968. This knowledge has been combined to develop a standard computer control package for concentrator control. Before the development of computer control package was started Outokumpu Oy had three large computer control systems. The first system was at the Pyhasalmi Concentrator, thee second at the LokkolaPlant, and the installation of the third system at the Vuonos Concentrator was just started.The characteristics of general computer software packages - both control and executive ones - were thus well-known as well as the practical usage of control theory. Furthermore Outokumpu had developed the Courier analyzer which included a so-called minicomputer. This project had given experience both 1n totally self made real tire executives and in computer hardware and peripheral d=vices. The instrumentation of the Outokumpu concentrators was designed and installed traditionally by the engineers of Outokumpu themselves."

Jan 1, 1973

By C. R. Edwards, J. R. Smith

"INTRODUCTIONThe purpose of this paper is to give an overview of the ALSANDS PROJECT, with emphasis on the approach taken in the design of the Primary Extraction and Froth Treatment Plants.The Alsands Consortium (Table 1) is a Joint Venture of 9 companies formed in 1978, with its prime objective to obtain governmental approval and concomitant suitable commercial terms to build and operate a third oil sands mining and upgrading facility in the Athabasca Oil Sands Deposit.Alsands Energy Limited is the Consortium Manager.The Alsands Project will provide about 10 percent of Canada's forecast crude oil production by 1990, which will be 20,700 cubic metres per day (130,000 barrels/day) of synthetic crude oil plus 1,100 cubic metres per day (7,000 barrels/day) of liquid petroleum gases. To construct Alsands the time required by construction labour, field non-manual and engineering support personnel as well as those trades involved in project-related urban development will total approximately 100,000,000 manhours or 50,000 man years. During regular operations, 4,800 people will be employed by Alsands.Approximately $13 billion will be spent to design, construct and start up the plant by the scheduled 1988 completion date."

Jan 1, 1982

By M. Y. Solar, W. G. Davenport

Most of the recent growth in nickel demand has been for the production of stainless steels and virtually all of it has been satisfied by new ferronickel and nickel pig iron projects. But "good" laterite resources are more and more difficult to find: Ni grades are decreasing while Fe contents are increasing and SiO2/MgO ratios vary from 1 to 3 and even higher if limonites are included. Predicting the smelting behaviour of such ores requires a more fundamental understanding of the thermodynamics of these systems in commercial operations. One valid question is thus whether these furnaces are at thermodynamic equilibrium or not. The present paper investigates this question, compares calculated and measured O2 partial pressures and mechanical/chemical slag loss ratios and concludes that ferronickel furnaces are not at thermodynamic equilibrium. This departure from expectation is actually to the operators’ advantage: the targeted ferronickel grades can be obtained at lower reduction potentials than calculated from pure thermodynamics.

Jan 1, 2014

By S. Kelebek, E. Yalcin

"Two composite samples of a gold-bearing ore from North Western Quebec (Barry Township) were tested to assess their grindability, recovery by flotation and Knelson concentrator as well as extraction using cyanidation.Work index values (BWi) for two composites were found to be 10.4 and 11.0 kWh/tonne, respectively. Flotation tests involved feed samples for which the particle size ranged from a P80 value of 205 ?m to 53 ?m, which resulted in gold recoveries of 91.8% to 95.8% (at gold grades of 130 to 140 g/tonne). Cyanidation tests indicated gold extraction levels ranging from 94.2 to 97.5%. Higher, recoveries were obtained at finer grind sizes and higher cyanide strengths. However, the grind size had a bigger impact on gold leaching. Gold grade and recoveries obtained with the Knelson concentrator device were lower than those obtained by flotation. Mass recoveries to concentrate products strongly depended on the fluidization pressure used. High gold recoveries can be obtained generally at low water pressures of 3 to 4 psi. Higher water pressures produced more selectivity at the expense of higher gold losses.It can be concluded that the gold ore samples tested have good characteristics due to moderate levels of energy requirement, high gold recoveries and extraction rates attainable.INTRODUCTIONBench scale testing of ore samples is an important step in the long series of metallurgical and economic studies involved in bringing new ore bodies into production (Coleman, 1980). Through a laboratory testing program, the behaviour of an ore in a particular process on a larger scale can be established. In an exploratory manner, this testing can be helpful for classification of ores in terms of “good ore” or “problematic ore” by providing preliminary information on recoverability of target minerals. Potential problems if any can be assessed by further investigation of products (e.g., middlings) obtained as a result of testing. A detailed bench scale experimental program in the metallurgical laboratory can also provide the background work for flowsheet development. Depending on the simplicity or complexity of the ore to be tested and type of details required from a testing program, the amount of samples required in the mineral processing laboratory can vary from several kg to more than one hundred kg depending on how detailed the metallurgical testing program is."

Jan 1, 2008

"In many gold deposits, a reasonable amount of the gold that remains in the cyanide leach tailing is associated with pyrite. Some gold plants have found it economical to incorporate gravity circuits on the cyanide leach tailing to recover the gold bearing pyrite before subjecting it to fine grinding, then recycling the ground pyrite back to the cyanide leach circuit. For low grade gold deposits the loss of gold bearing pyrite is exacerbated as the broader process economics normally favour a coarse grind (> 150 ?m). An option that has received attention over the last decade is to recover the gold bearing sulphides in the milling circuit by gravity, similar to the widely-practiced gravity recovery of free liberated gold. The modelling and simulation of a gravity gold circuit in a milling flowsheet is relatively simple compared with a gold bearing pyrite circuit, as the recycle advantages and breakage functions for gold and pyrite are quite different. This current work represents an attempt to simulate the gravity recovery of gold bearing sulphides in a milling circuit that also incorporates fine grinding, using relatively simple simulation models. The simulation results are further explored with reference to the economics of this type of circuit.INTRODUCTION It is well known that gold is associated with pyrite in most sulphide gold deposits and that the amount of gold associated with pyrite will vary from deposit to deposit. Most of the gold associated with pyrite is liberated in the grinding circuit during the processing of a gold ore. However, some gold remains occluded in the pyrite and is lost in the final leach tailings. Little information is available in the literature on methods to enhance gold liberation from pyrite by specifically removing pyrite, either in the grinding circuit or from the final tailings stream, and grinding it finer. One of the earliest record of this approach is described by King (King, 1949) where pyrite present in Witwatersrand gold ores was recovered by circular jigs and Johnson barrels followed by fine grinding of the concentrate prior to cyanide leaching. Dominy et al., 1970, provides information on the use of Yuba jigs to recover pyrite from the discharge end of ball mills. Again, the jig concentrate was ground to a fine particle size before cyanide leaching to recover the liberated gold. Unfortunately, no information is provided in these articles on the additional gold recovery from implementing these circuits in a flowsheet. Martins et al., 1993, provides details of a gravity circuit using Reichert cones and spirals to recover gold bearing pyrite from a cyanide leach tailings followed by fine grinding of the pyrite concentrate in a Tower mill. The ground concentrate returned to the front end of the already installed leach circuit. Overall recoveries were increased by some 2% to 4% depending on the amount of gold and pyrite in the plant feed."

Jan 1, 2017

By Jochen Franke, Derek D. Lichti

"This paper gives a brief introduction and presents recent enhancements in the MillMapper technology developed for thickness gauging and condition monitoring of liners in grinding mills used for mineral comminution. It also introduces the recent expansion of MillMapper functionality aimed at mill operational optimization. MillMapper uses calibrated laser scanner data captured inside a mill to determine thickness at several million points on all liner surfaces at feed, shell, and discharge ends with millimeter accuracy. Data capture takes nominally 5 minutes, hence interruption to mill operation is negligible.The paper is structured as follows. First, a review of the technology and discussion of recentlyimplemented innovations such as liner change out forecasting, automatic ball charge volume and weight measurement, and automatic mill liner volume/weight measurement for mill operational control or load cell calibration are presented. Combined weight information provides control over load related issues such as load fill levels for relining, or start-up load after reline. Feature descriptions are followed by a listing of the numerous advantages over existing techniques and the benefits to operations. Results from the application of MillMapper are presented throughout.INTRODUCTIONThe need for thickness gauging of liners of grinding mills used for mineral comminution is well known but the task is rarely completed to a satisfactory standard. Liner failure resulting from excessive wear can have expensive consequences. Furthermore, mill performance as well as liner wear are known to be correlated to lifter bar geometry (Napier-Munn et al., 1999; Rajamanni et al., 2005). At present, the following techniques are routinely used to gauge the thickness of mill liners:1. Visual inspection;2. Ultrasonic thickness gauging (for steel liners); and3. Physical inspection (e.g. nail inserted in rubber liners)"

Jan 1, 2008