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By Glenn Clarke, Ben Holloway, Barry Lumsden

"BHP-Billiton’s Cannington mine in Northwest Queensland produces a silver-rich lead concentrate and a zinc concentrate. The plant has the capacity to treat 3.1 mta and for the year to June 2006 mean feed grade was 10.3% Pb, 461 g/t Ag and 3.7% Zn. Recovery losses of lead and silver are primarily in the fine fractions. In 2001, the plant was modified when it was determined that splitting the lead feed into two size fractions, and floating the two fractions separately increased fine lead and silver recovery. The fine lead fraction (called the split float) is 80% <16µm and about a quarter of the lead and silver is in the finest fractions. In 2006 to further improve lead and silver recovery a statistical plant trial of magnetic aggregation conditioning of the split float feed was undertaken. While pure galena is diamagnetic the Cannington galena contains iron within the galena mineral resulting in the galena having a paramagnetic, magnetic susceptibility. The trial demonstrated to greater than 98% statistical significance an increase in lead recovery, and to 99% statistical significance an increase in silver recovery, and to greater than 97% significance an increase in the lead in concentrate in the split float circuit. These economic recovery benefits with magnetic aggregation of the split float feed, has led to the incorporation of this technology in the plant.INTRODUCTIONThe BHP-Billiton owned Cannington silver, lead and zinc mine is located in the Australian state of Queensland about 800 kilometres west of the coastal city of Townsville. This is a pastoral area characterised by large cattle farms, low population density and very little infrastructure. The area has two seasons, a wet, hot summer and a dry, mild winter. Because of its isolation and lack of infrastructure the mine has always operated on a fly in fly out basis.Geological drilling defined the resource as 43.8 million tonnes grading 11.6% lead, 4.4% zinc and 538 ppm silver (Walters and Bailey, 1998). At times during its history Cannington has held the crown as the single largest lead and silver mine in the world (Torrisi and Smith 2003). Another unique characteristic of the operation’s history is the rapid progress from discovery to operation. The first drill to intersect the ore body was in 1990 and the mine and plant was commissioned in 1997."

Jan 1, 2008

By J. M. Hemstock, H. L. Ames

THESE forums have been held annually since they were introduced by Mr. L. E. Djingheuzian in 1957. The response of millmen to questionnaires such as the one this year is ample evidence of the interest these create. This year's topic is unlike those of past years since it is an attempt to understand a new concept of grinding rather than to improve the old. Before, we were thinking in terms of 2-5 cents improvement in grinding costs, now we will be discussing savings of 15-2 5 cents. As a background for this forum we will begin by summarizing the major controversial issues relating to grinding with steel, many of which will undoubtedly apply to pebble milling. Our •task to-day is to learn what has already been done in pebble milling and obtain a fair overall view of this method of grinding.

Jan 1, 1961

By P. K. Gummer

Gold was fir t di covered on the Seabee property in 1947 by the Con olidated Mining and Smelting Company Limited. Acquired by Claude Resource Inc. in 1984, the property is now being explored by Placer Development Ltd . in joint venture with Claude Re ources. Between Augu t, 1985 and March, 1986, Placer completed 54 drill holes, extending the main o. 2 and o. 5 vein y tern to a depth of 1000 feet below surface and di covering two new satellitic gold zone 800 feet west of the o. 5 gold zone. Drill proven and probable geological re erve in four zone ( os. 2, 5, II, 14) to a vertical depth of 1000 feet, are approximately 1,120.000 tons averaging 0.32 oz. Au/ton.

Jan 1, 1986

By P. J. Sylvester, D. H. C. Wilton, R. I. Wheeler, M. A. E. Huminick, D. Evans-Lamswood

"Abstract — Systematic characterization of the mineralogy of massive sulfide ore deposits can provide valuable insights into ore genesis and metallurgical processing that would not be apparent otherwise. We describe a method wherein (1) normative mineral abundances are calculated from whole-rock chemical assays using an algorithm, (2) the results are cast in block models, (3) the results are then verified with automated image analysis of a subset of samples prepared as grain mounts by backscattered electron imaging and X-ray mapping on a scanning electron microscope, and (4) those verified results are integrated with textural data determined by optical microscopy.The technique is applied to the magmatic Ni-Cu-Co sulfides of the Ovoid orebody, based on geochemical assays of 3175 whole-rock samples. Three principal ore zones are defined in the Ovoid. Type I ore occurs in the center of the deposit; it is magnetite-rich and pyrrhotite-poor with intermediate contents of pentlandite and chalcopyrite. Type II ore surrounds Type I ore and is pentlanditechalcopyrite- rich with intermediate contents of pyrrhotite and magnetite. Type III ore occurs at the base and along the outer periphery of the northern part of the orebody; it is pyrrhotite-rich and chalcopyrite- poor with an intermediate content of pentlandite, and a low-to-intermediate content of magnetite."

Jan 1, 2012

By Marcel M. Djivre

Electromagnetic wire rope testing is the only practical non-destructive method to evaluate the current condition of wire ropes in general, and in the mining industry in particular. The method has been successfully used worldwide for decades. In mining it has considerably improved the safety of operations. Yet, problems persist. In particular both the Loss-of-Metallic-Area (LMA) readings, as well as the Local Fault (LF) readings, as obtained by different instrument operators, were found to be unacceptably diverse. Consequently, large sums have been spent internationally, as well as in Canada, comparing a whole range of these instruments in an effort to rate their performance. The published results clearly confirmed operational experience, namely that the LMA and LF readings were, indeed, unacceptably diverse. Excellent pertinent conclusions were also published. In Canada the authors? conclusion was that while the divergence in the LF signal based results could, possibly, be attributed to both hardware and human-error related factors, the LMA signal related quantitative uncertainties were human-error related. Consequently, much attention was paid in Canada to developing novel control methods, namely computerized ones to replace the more error-prone analogue ones. In addition new sensor-heads were developed, both for portable as well as for permanent installation with remote control operations. While these efforts were of considerable benefit, the situation must again be carefully reviewed with very special reference to deep-level mining, now of immediate concern in Canada as well. In this paper the authors supply an overview of the results obtained by three of the four major international projects referred to previously. Relevant conclusions/comments are also restated. Moreover, specific recommendations are put forward as to how current problems might best be remedied.

May 1, 2004

By Jessica Egan, Daniel Hodouin, Mohammed Reza Khalezi, Steve Bellec, Claude Bazin, Carl Duchesne

"A model for gold leaching was developed and implemented into a simulator. The simulator uses a description of the ore that is based on the mineral composition and precious metal content. The model accounts for the distribution of minerals and precious metals through the size intervals. An ore texture model based on precious metals and minerals grain sizes within the ore is used to predict the exposed surface of the species in the various size intervals. A model is proposed to predict the rate of dissolution of any species dissolution using the exposed surface, particle size, cyanide and oxygen concentrations, temperature, pH and agitation. The flow of slurry through a leaching tank is modelled using a parallel arrangement of perfect mixers to simulate a bypass and a Cholette-Cloutier mixer model. The simulator can be used for circuit optimization but was developed to provide an environment for learning the basic concepts of process analysis methods. Examples are given to illustrate the application of the simulator for planning tracer tests and sampling campaigns.INTRODUCTIONThe constant pressure to improve the operation of processes that are already close to optimum imposes to plant engineers to be aware of the available process analysis tools and to be able to use them to assess process operation and make a diagnosis to possible problems. Some of these tools are:•,Sampling the streams of a process;•,Data reconciliation;•,Analysis of mixing in reactors;•,Estimation of rate parameters;•,Optimal tuning and flow sheet modification;•,Automatic control and real time optimization.Although some of these tools are part of the curriculum of many university engineering programs, the students are not exposed to actual process analysis problems. In addition the complexity of the ore dressing processes compounded with unavoidable measurement errors and plant disturbances can discourage many motivated engineers. These factors often lead to an incomplete valorization of the collected data and a poor return on the investment associated with the sampling, testing and analysis of the samples.Simulators are readily available for process optimization (Yalcin and Shang, 2006) by allowing:•,The simulation of various flow sheet configurations;•,The simulation of a change of operating conditions;•,The simulation of various control strategies;•,The training of plant operators.However few simulators have been developed with the objective of providing an environment for training plant engineers and technicians to the use of tools for process analysis and optimisation."

Jan 1, 2008

By Gordan Marrs, O. Matwijenko

"The Kidd Creek metallurgical site has been highly integrated to maximize revenues from its base metals mines. The site's operations include a Concentrator processing 4.3 million tonnes per year of copper-zinc ore, a Copper complex producing 100,000 tonnes per year copper cathode and a Zinc Plant producing 130,000 tonnes per year of zinc metal. To improve operating efficiency with increasing emphasis on productivity, process control, process automation and process optimization have been augmented and expanded during the last decade.New microprocessor based control systems provided the opportunity for the capturing of an expanded number of process related measurements. As supervisory computers were already in place, databases could be generated with only increased disc space and database management programs required. Once the data had been accumulated and stored, a method was then required for handling, screening and sorting large volumes of information in relatively quick and easy terms in order that the stored information could be used. By 1984, a decision to use personal computers for accounting, engineering and management reporting provided the vehicle necessary to handle this data. Using commercially available software, large volumes of stored process data can be downloaded from existing databases, filtered, sorted, analyzed, graphed, displayed, projected, reported and distributed. All this can be performed in very little time."

Jan 1, 1989

By Samantha Beeson, Collin Poole

"This paper examines the development of two new selective flocculation techniques for recovering KCI values from low grade carnallite potash ore.Direct recovery of KCI was achieved using one of two mid-range molecular weight polymers, one non-ionic and the other strongly cationic, in conjunction with the same selective anionic dispersant. From initial response studies on pure minerals and selective flocculation tests on synthetic mixtures, investigations were extended to recover KCI values from a natural carnallite potash ore. A single-stage process achieved products grading up to 61.3% KCI at recoveries up to 85.4% for the cationic flocculant suite and up to 64.8% KCI with recoveries as high as 90.3% for the non-ionic flocculant suite, with dispersant and flocculant dosages of 150-250gr1 and 10-75gr1, respectively. These results clearly demonstrate not only the feasibility of treating such problematic ores at low reagent dosages, but also the potential elimination of desliming prior to froth flotation.The experimental results presented are considered in terms of the reagent adsorption phenomena present in the system and various aspects of floe formation and cleaning.IntroductionAs high-grade sylvinite potash ores become scarce, lower grade sylvinite and carnallite potash ores are becoming more attractive. However, these ores often contain large amounts (up to 15% by weight) of water-insoluble slimes which must be removed prior to KC! flotation, owing to their high adsorptive capacity for the amine collectors used. The slimes are generally removed by scrubbing followed by hydroclassification and any residual material not eliminated is blinded with reagents such as guar gum or starch to inhibit interference in subsequent KC1 flotation.An alternative approach used commercially in Canada by Agrium Inc. involves water-insoluble slimes flotation prior to KC! flotation. In this process, the slimes are selectively flocculated and then separated by froth flotation. Losses of KC! to the slimes product and associated brine requirements are both reported to be less than those encountered in conventional desliming, but the technique is somewhat limited in its application at present."

Jan 1, 1997

By T. S. Heah, S. T. Bishop

"The Lorraine property is located in the northern Intermontane Belt, north-central British Columbia. The property is underlain by intrusive rocks of the Duckling Creek Syenite Complex, an Early Jurassic, alkaline phase of the Late Triassic-Early Cretaceous Hogem Batholith. Two significant zones of copper-gold mineralization have been explored to date; the Main zone (historical Upper and Lower deposits) and the recently discovered Extension zone. A combined geological resource was estimated in 1975 for the Main zone deposits at 10 million tonnes averaging 0.7% Cu, and from 0.1 g/t Au to 0.34 g/t Au. The Extension zone, still at an exploration stage, is estimated to be considerably smaller than the Main zone.Copper-gold mineralization at Lorraine Mountain exhibits characteristics suggestive of both magmatic and hydrothermal origins, and may be related to orthomagmatic-hydrothermal fluid flow contemporaneous with magmatism and development of migmatitic fabrics. A depth greater than the depth of formation of classical porphyries may explain many of the variations displayed by mineralization at Lorraine. These characteristics distinguish Lorraine from other members of the alkalic suite of porphyry Cu-Au deposits in the Canadian Cordillera.IntroductionThe Lorraine prospect is located in the Omineca Mountains, at the headwaters of Duckling Creek, 56 km northwest of Germansen Landing (Fig. 1). The property is centred at latitude 55°55'N, longitude l25°26'W on NTS mapsheet 93N/14W. The area is typified by mountains of moderate relief, with elevations ranging from 1150 m in the valleys to peaks of 2000 m. Valleys are U-shaped and blanketed by glacial till which give way to steep, talus covered slopes and sharp ridges. Outcrop is generally limited to elevations above 1600 m (Fig. 2)."

Jan 1, 1995

By Kevan J. R. Ford, Chris A. Fleming, Robert D. Henderson

"Kinross Gold Corporation’s Maricunga mine in Chile presents an ideal opportunity to apply the SART process to a heap leach operation. Given increasing levels of cyanide complexed copper in current pregnant solutions from Maricunga heaps (over 300 ppm Cu), there is a need to reduce these levels and decrease cyanide consumption and cost. Additionally, a means of disposing of this copper is required, preferably as a saleable product. The well published SART process applied to gold-copper cyanide solutions was selected as it allows rapid precipitation of cyanide soluble copper as a relatively pure copper sulphide. The process also allows effective recovery and recycle of cyanide back to the heap leach operation.This paper reviews the SART process chemistry and presents a summary of SART bench-scale and pilot plant testwork completed by Kinross. Key design parameters and engineering aspects of the proposed plant at Maricunga are discussed. SART plant operating costs and net revenue analysis are included and illustrate the compelling economics of installing a SART process plant. INTRODUCTIONMaricunga OperationCompañia Minera Maricunga (CMM), fully owned by Kinross Gold Corporation, operates the Maricunga mine (formerly Refugio) which is located 120 km due east of the city of Copiapó at an elevation of 4,200m above sea level in the Andes mountains of northern Chile. Figure 1 shows a map of the location of the Refugio mine relative to other gold and copper mines in the region."

Jan 1, 2008

By S. B. Mishra

Ni-20Cr metallic coating was deposited on a Ni-based Superalloy (3Fe-19.5Cr¬0.3Ti-Balance Ni) by a shrouded plasma spray process; NiCrA1Y was used as a bond coat. Erosion studies were conducted on uncoated and plasma-spray coated superalloy specimens using an air-jet erosion test rig at a velocity of 40 m/s and impingement angles of 30° and 90°. The coating has been characterized for porosity, microhardness and microstructure and X-Ray Diffraction (XRD). An electron probe micro analyzer (EPMA) was used to obtain X-ray elemental mappings of the coating along the cross-section. Scanning electron microscope (SEM) with EDAX attachment was used to analyze the eroded surfaces.

Jan 1, 2005

By L. K. Sinclair

In situ leaching of copper has the potential to deliver low-cost recovery for soluble copper deposits of sufficient permeability. In order to provide key inputs for modeling of the in situ leaching process, bench scale test work can be used in conjunction with geochemical software to identify key leaching and precipitation reactions. A case study is presented in which a series of flow-through acid leaching tests were conducted on drill core sections from a porphyry copper deposit without crushing or grinding. Concentrations of copper and other key elements were monitored in the recirculating leach solution over several months. A geochemical model of each leaching test was prepared using Geochemist&apos;s Workbench using a combination of diffusion-controlled and reaction-controlled rate mechanisms. Key reactions included leaching of chrysocolla, iron oxides, and plagioclase. The model was able to simultaneously satisfy the evolving chemistry of the recirculating leach solution, mineralogy of the head and leach residue, measured acid consumption, and leaching rates for key minerals reported from previous laboratory studies. Opportunities and limitations for scaling up the model to simulate the commercial wellfield are discussed.

Jan 1, 2015

By Hiroaki Sakurai

Using batch reactors equipped with CO and CO2 gas sensor, CO oxidation over Au catalysts in room air has been investigated and compared with Pt catalysts. For both Au and Pt catalysts, the logarithm plot of CO concentration data during initial 10 min are well fitted by straight line, which mean the first order dependence on the CO concentration between 10 and 10,000ppm. The rate constant value (k) over Au and Pt catalyst generally decrease with increasing initial CO concentration above 100ppm. The value decrease especially rapidly between 100ppm and 1,000ppm over Pt catalysts, which should be due to the strong adsorption of CO on Pt. The k value over Au catalyst has become ten times or more as large as the Pt catalysts at 1,000ppm. Already for the initial concentration of 100ppm CO, the strongly adsorbed CO on Pt metal has been clearly observed by FT-IR during the reaction over Pt/Al2O3. A part of the strongly adsorbed CO has been remained on Pt after the disappearance of gas phase CO, and has gradually changed to gas phase CO2. Such strongly adsorbed species on Au has never been observed on Au/TiO2 under the same reaction conditions. These features of Au catalysts are favorable for indoor air purification.

Oct 1, 2003

By L. E. Djingheuzian

Introduction Cylindrical grinding mills have been used for a long time, but during many years of testing no methods were established that gave consistent results. The work by both scientific investigators and operators led very often to contradictory conclusions, and each millman was very much &apos;on his own&apos; when trying to find the limits for working of his ball-mills. Fred C. Bond, in the introduction to his paper on Wear and Size Distribution of Grinding Balls, freely states that the mathematical immaturity of the grinding process "is the cause of many inefficient designs and installations as well as of much poorly oriented experimental work", and evidently he sees the solution of all our grinding problems in the discovery of the fundamental laws "which can change the empirical art into a mathematically exact science". While the writer is in complete agreement with Mr. Bond that grinding is governed by fundamental laws which can be expressed by mathematical formulae, there are so many variables in grinding, all interwoven with each other and mutually influencing each other, that the coalescing and merging of all these variables into a harmonious mathematical whole appears to be a homeric job. However, a light appeared on the horizon when it was proposed that power be used as an index of the grinding operation. It is pertinent to quote Gow, Guggenheim, Campbell, and Coghill from their paper on Ball Milling, in which they say: "Much of the past literature devoted to the subject has been concerned with the Kick-Rittinger dispute over evaluating grinding, or with dissertations on ball paths and the closely allied impact versus attrition hypotheses or with the exclusive consideration of certain specific variables, such as mill speed or circulating load, irrespectve of the other factors in operation. Such subjects may be dismissed with but little discussion in what may be called, by comparison a new philosophy in the stud of ball milling. Briefly this &apos;new philosophy&apos; consists &apos; of using power as an index of the mill operation, and investigating grinding from the standpoint of set and induced variables with emphasis on the latter". But it is safe to assume that it was the work by E.W. Davis, H. E. T. Haultain, F. C. Dyer, A: M. Gaudin, J. Gross, S. R. Zimmerley, and others, that cleared the path for the authors of Ball Milling and induced them and A. W. Fahrenwald to look farther afield and thus to establish a far-reaching relationship between the power and capacity and the diameter of the ball-mills.

Jan 1, 1949

By G. E. Viens, R. A. Campbell, G. N. Banks

A survey of the literature reveals wide differences in the procedures used in making iron ore pellets and in the properties of the pellets obtained. In this paper, the methods of production and the properties of pellets, as reported in the literature, are examined in an effort to determine what correlations exist between them. An attempt is also made to determine the levels of the properties of pellets which are considered acceptable and desirable in modern practice. M UCH has been written regarding the important changes which have been taking place recently in the iron ore industry. As the demand for ore has increased and the supply of direct-shipping ore has dwindled, tl1e problems associated with the utilization of lower grade ores ha\•e been intensely investigated and important advances in both beneficiation and agglomeration techniques have taken place. In particular, there has been a great deal of activity in connection with the pelletizing method of agglomeration. The superiority of pellets over raw ore has been demonstrated in blast furnace tests

Jan 1, 1962

By H. D. Stolovitch, M. L. Lavallée

"An engineering system used to analyze and improve the performance of industrial grinding circuits has been adapted into a training program. The systematic approach to the design of training was applied in order to:Determine if training was an appropriate and effective tool for the transfer of this new technology;Define the technical content of the program;Establish program specifications based on learner and context characteristics;Design and test prototypes of the instructional modules; and,Establish a strategy for plant implementation and maintenance of the training program.This paper thus describes how a engineering system is transferred to the workplace through the use of designed instruction.INTRODUCTIONTraining can change human performance in a permanent way. It can improve productivity for the overall benefit of an operation. Organizations can therefore use training as a mean to achieve their specific goals. Training is not the solution to every performance problem. However, in many cases, it can be a very effective way to attain desired performance.The benefits of training are often difficult to assess. However, to remain competitive, organizations must introduce new products and innovative technologies. Often, they can do both most effectively through suitable training interventions.In order for the effort and investment of training to provide optimal results, two points of view should be adopted.• An overall view of the system in which training is to be applied.• A systematic approach to the design of training (Stolovitch and Keeps, 1989)."

Jan 1, 1992

By Colin Hardie, Erik Bartsch, Guy Comeau

"The Xstrata Nickel Raglan Concentrator, located at the northern limit of Quebec’s Nunavik region has been in operation since 1997. Originally designed to process 800,000 tonnes per year of high grade nickel ore, plant capacity is currently approaching 1,200,000 tonnes per year due to an Optimization Project (FAG to SAG conversion) in 2005 and many recent processimprovements.In 2006, the Raglan Metallurgical Group with assistance from the Xstrata Process Support group decided to make SAG control one of its priorities in order to further increase throughput and improve stability. The original primary mill control logic delivered by the Optimisation Project in 2005 consisted of a simple PID loop controlling the mill charge (bearing pressure) through feed rate changes. Due to the multivariable, non-linear and disturbance dominated nature of the SAG circuit it was quickly found that this traditional PID control strategy was not able to adequately control the mill; resulting in poor stability, frequent operator interventions and less than optimum performance.This paper describes the successful integration of advanced field systems such as mill feed image analysis (Wipfrag) and crusher gap controller (ASRi), into a multi-variable fuzzy logic SAG mill controller. The process of how a strategy for control was developed and implemented directly in the existing control system (exploiting standard system capabilities) is discussed. It is also shown how this approach has provided mill automation which is simple, robust (despite erratic feed characteristics), and delivers both an increase in throughput and a reduction of variability.INTRODUCTIONOriginally commissioned as a Fully Autogenous Grinding (FAG) circuit in 1997, preparations to convert the Katinniq Concentrator Mill to SAG had in fact commenced as early as 1999, but were halted owing to alternative process improvements which achieved up to 1Mtpa as early as 2000. Up until this time, plant process controls had remained more or less static; modifications having been limited to those required by new equipment. Primary mill process controls and related measurements were at this stage limited to:"

Jan 1, 2008

By Yannis Faitakis, Simon Bergeron

"This paper describes the implementation of model-based predictive control (MPC) on a vacuum heat exchange network at Alcan's Jonquière Complex. The plant produces alumina from bauxite ore using the Bayer Process.The MPC system was commissioned in 2000, by Alcan personnel and an Advanced Control Engineer from Matrikon. The application has achieved an uptime of about 97%.The benefits were:• Energy savings : 0.5% - live steam reduction• Reduced variation: standard deviation of key temperatures reduced by 50%• Plant throughput increase: via the automated use of surge tanks• Decreased supervision time: by plant operations• Faster and easier diagnosis of problems and plant constraintsThe benefits were higher than expected with a payback time around 12 months.The purpose of the vacuum heat exchange network is to recover energy from a hot stream (pregnant liquor) by transferring it to a cold stream (spent liquor). The system is very tightly coupled, and the main constraints of the multivariable control system are: • to control the many tank levels involved in the process (8 tanks)• to maximize the temperature of the heated streams (spent liquor)• to prevent cooled streams (pregnant liquor) from cooling to below a minimum temperatureThe system includes 9 manipulated, 14 controlled and 2 feedforward variables.The advanced control implementation runs under an expert system environment and provides added flexibility to the overall control strategy."

Jan 1, 2003

By Ken Norris

"Campbell Mine which has been in continuous operation since 1949 has undergone numerous operational changes in the past 3 years. This paper describes the process changes which have occurred at the Campbell Mill since 1991. The paper discusses the engineering, construction and startup of the effluent treatment circuit, the new oxide filtering and merrill-crowe precipitation circuit, and the changes to the pressure oxidation circuit. Particular attention is paid to startup problems and solutions to these problems. A review of the benefits and results of these process changes is also discussed. Future changes are also briefly reviewed with anticipated goals and expectations.IntroductionCampbell Mine is a gold mine 100% owned and operated by Placer Dome Canada. It is located in Balmertown, Ontario which is located 300 kilometres northeast of Winnipeg, Manitoba. The Campbell Mine has been in continuous operation since 1949, producing nearly 8,000,000 ounces. In 1993 Campbell Mine is budgeted to produce 289, 750 ounces of gold at a cash cost of $157 US/ounce. Daily mill tonnage is budgeted for 1,430 tons/day but has milled as high as 1,600 tons/day on a consistent basis. Recovery in 1993 is 93.4% and is budgeted to average 94.5% in 1994.In the 1992 edition of the proceedings of the Canadian Mineral Processors conference I described the changes that were made at Campbell starting in 1989. This paper continues where the 1992 paper ended and discusses the changes to the Campbell mill since that lime. The changes to Campbell since January 1992 have been numerous with capital expenditure in the mill exceeding $14.1 million. This paper describes three primary projects since that lime: the effluent treatment project ($5.5 million), the oxide leach/filtering and merrill-crowe project ($4.6 million) and the pretreatment project ($2.81 million). The focus of this paper is to highlight the justification of each project, the design and engineering, and some of the initial operational problems and solutions applied to those problems."

Jan 1, 1994

By Jr. Blanchette, Claude Bazin, Paul Blatter

"Reporting of the gold production during a month implies to estimate the mass of gold that is processed by a mill, recovered as dorés, lost to the tailings and the change of gold inventory within the plant equipments. Carbon in pulp and/or carbon in leach tanks may contain gold inventory as large as 72 000 g (2 200 troy oz) of gold which needs to be estimated from measurements obtained from sampling the material contained in a tank. Sampling and analysis introduce measurement errors that could be systematic and/or random. This paper examines the effect of measurement errors on the estimation of the gold inventory and on the variation of the inventory in the carbon in leach tanks of the Agnico-Eagle Mines Limited Lapa concentrator. Results of the analysis demonstrate that the inventory is measured with ±10% reproducibility, while the variation of inventory used for mass balance is known with ±300% reproducibility.INTRODUCTIONThe preparation of a production report implies to reconcile measured data with the constraint of material balance over the considered period of production. The sampling and data reconciliation problems were recently reviewed by Morrison (2008) and Jansen (2009a; 2009b). Typically a production report is prepared to cover a period over which the mass conservation hypothesis can be considered valid. In gold processing plants, the change of inventory can seldom be neglected and plays a significant role in material balance calculations. This is particularly true for circuits using carbon adsorption, where the complete cycle of carbon is measured in days.In a 2000 t/d gold ore processing plant, the inventory can be as large as 140 000 g of gold in gravity concentrator bins, carbon in pulp tanks and in the electrolysis circuit. This is equivalent to the mass of gold that fed the circuit during 10 days of operation with 7 g/t feed assays. The change of inventory may therefore strongly impact the production figures and especially the plant recovery at the end of a week or a month. Despite the financial impact associated with gold inventory, few publications are devoted to the subject. Bruington (1991) and Morrison (2008) discussed the problem and stressed the need for the development of reliable methods to estimate the gold inventory in carbon in pulp or carbon in leach circuits."

Jan 1, 2011