Search Documents

By R. F. Pilgrim

"This first chapter of the SPOC Manual 1s intended as a general introduction to all the other chapters in which the mineral process evaluation and optimization methodology is presented. This chapter also provides a brief description of available software for mainframe and m1crocomputers and provides information on its distribution by the Technology Information Division. CANMETRESUMECe premier chapitre du manuel SPOC est un guide général qui présente les autres chapitres ou l'on traite de la méthodologie d'évaluation et d'optimisation des procédés miméralurgiques. Le lecteur y trouvera aussi une description sommaire des programmes de calcul développés pour ordinateurs centraux et pour micro-ordinateurs ainsi que la façon de les obtenir de la Division de l'information technologique du CANMETComputer simulation of mineral processes has slowly emerged as an established branch of activity in A&D centres and in several teaching institutions after decades of slow progress. Substantial efforts to develop application software based on years of research on process modelling are being pursued in many countries ( 1 .2,3,4.5). The late BO's should see a flurry of success stones about significant leaps of productivity achieved in several plants through systematic process evaluation and computer simulation studies.In Canada. the SPOC project (Simulated Processrng of Ore and Coal) was initiated at CAN MET in 1980 with the obJective of developing and transferrrng to the mineral rndustry a computer-based methodology for process evaluation and optimization. The proiect deliverables include:- an engineering manual covering most aspects of off-line computer simulation:- a bank of FORTRAN software accessible through data communication lines for user familiarization and distributed at a nominal cost in source form by CANMETThe yearly progress reports of the project contain detailed information on the project elements since 1980 16). The context of SPOC can be summed up by quoting Gordon Agar""s 1975 statement during a workshop on research needs in the mineral industry: ""What is needed is a joint effort between the academic community and one or more enliglitened industrial concerns. The effort will have to be an iterative one with data collection followed by testing the model modification of the model (s), more testing prediction, modification. data collection and so on (7). A few parallel in1- t1atives had been reported in papers or research contract reports (8, 9) but there was an obvious need for CANMET to bridge the gap between the R&D and the production sectors by undertaking a project that would document past applications and induce a new generation of initiatives through the consol1dat1on and transfer of suitable manuals and software. The effort took shape as the CANMET SPOC pro1ect in which methodology transfer was to play a key role (10)."

Jan 1, 1986

By T. K. Dichmann, J. A. Finch

"Microflotation, laboratory batch, and in-plant minicell continuous tests were conducted to determine if copper ions not only activate sphalerite but also depress pyrite when the minerals are together. Microflotation tests showed that pyrite recovery increased with addition of copper ions in the pH range of 5-10 when alone, but decreased when sphalerite was also present. Batch flotation tests (Red Dog ore) showed that upon addition of copper, pyrite recovery decreased significantly. The continuous tests (Louvicourt Mine) confirmed this and indicated that with excess copper pyrite recovery can increase. The results suggest there is competition for copper and xanthate in favour of sphalerite. To take advantage of this competition it may be necessary to control the sequence and rate of copper, collector, and lime addition. INTRODUCTIONFlotation, in common with all physical separations, is an imperfect process that results in misplacement of minerals. One source of this imperfection is accidental activation by contaminating metal ions. An important practical example is the misplacement of pyrite to zinc concentrates. Flotation of sphalerite generally requires activation by copper ions, added as copper sulphate. Copper ions are also used to activate pyrite [l] hence there is the suspicion that accidental activation contributes to the recovery of pyrite (Py) with the sphalerite (Sp). However, recent literature has shown that copper ions may make Py less floatable if Sp is also present. Xu et al. [2] conducted continuous minicell tests in-plant at Kidd Creek Operations (division of Falconbridge Ltd.) and showed that while the addition of copper sulphate improved zinc/iron selectivity, it did so as much by reducing the recovery of pyrite as increasing zinc recovery (Figure 1)."

Jan 1, 2000

By Tatiana Ravani von Ow, Leandro Bomvisinho

"The size of mills driven by ring-gear or by a gearless solution has significantly increased in the last years. This has created additional challenges for the driven equipment. The requirements for the mechanical devices and the power grid have become more demanding. Controlled soft starting and operation, especially in dual pinion mills, as well as the protection of the equipment has gained importance. The selection of the drive solution for the mill therefore becomes more critical. Besides the classical criteria such as performance, efficiency and capital costs other factors come in the evaluation of the system such as flexibility and optimization of operation, reliability and the ageing of the grinding equipment.Variable speed solutions offer significant advantages over conventional fixed speed drives due to their application specific features and process optimization capabilities, even for ball mills where these are not seen yet as a must.The initial capital costs are higher than some of the direct on-line solutions however the benefits for the operation during the lifetime of the mill generated by the use of converter solutions justify the additional higher investment.INTRODUCTIONThe method of powering mills such as AG, SAG and Ball has a long and technically interesting history. It can be said; almost every decade in the last half century had its own trend. The trend was dominated by the market demand, by the process development and requirement of the mechanical and the electrical technology but mostly by the price of the applied technology.All has started about 50 years ago with mill diameters less than 20 feet driven by fix speed synchronous or wound rotor motors with the power of about 2 MW. Around 10 years later, the mill sizes increased, but the necessary torque could not be transmitted by one pinion therefore dual pinion solutions were developed to power the mills. However, realization of the dual pinion solution was only possible when load could be evenly shared between the two pinions."

Jan 1, 2010

By S. Banisi, M Nesari, J. A. Finch, N. Nobari, H. Javani

"The flotation circuit of the Sarcheshmeh copper mine, which processes 40,000 t/d of ore at 0.9% Cu, consists of rougher, cleaning and recleaning stages. Mechanical flotation cells are used in all stages. To investigate the amenability of flotation columns in the plant especially for cleaning and recleaning stages, a column was designed, constructed, and installed at the Sarcheshmeh pilot plant. The pilot plant, which has an identical circuit configuration compared to the plant, can process up to 2 t/h of ore. The overall rate constant of the ore along with the target recovery of 94% were used to size the column. A 26 cm by 550 cm column with two filter cloth spargers and a submerged froth washing system using perforated tubes was constructed. To detect the interface level a 180 cm conductivity probe consisting of 1.5 cm diameter ring electrodes was installed. The tailings flowrate controlled by a variable speed pump was used as the main control variable. The initial testing program revealed that the optimum superficial gas and wash water flowrate were 1.5 and 0.3 cds, respectively. It was observed that the cleaning stage grade when using the column could increase by 2-3% at constant overall copper recovery.IntroductionThe Sarcheshmeh copper ore body which may rank as third or fourth largest in the world contains 1 billion tomes averaging 0.90 % copper and 0.03 % molybdenum. It is in located south-east Iran and has been processing 40,000 t/d of ore since 1982. After three stages of crushing, the ore with a F80 of 12.7 mm is fed to 8 parallel ball mills in closed circuit with cyclones to produce a product 70% finer than 75 microns. The concentrate of the first stage of flotation is reground and the tailings constitutes the main component of the plant final tailings. A concentrate with an average grade of 31% Cu (chalcopyrite and chalcocite) is obtained after cleaning and re-cleaning stages in mechanical cells. Total recovery of the plant varies between 83-87 % depending on ore type and operating conditions."

Jan 1, 2000

By A. Rajan

"Drawing upon the experiences of a gold extraction plant startup in the Southern Pacific, this paper will describe the importance of the commissioning phases to the project outcomes to achieve early target production despite an overall time compression to attain full plant production. Some of the highlights of the paper include: impact of roles played by the Owner’s construction and operations representatives during project commissioning, the formation of a capable and competent “Mod Squad,” availability of trained operating personnel, control over work priorities and the scope of work needed to bring the plant to safe initial operations. The commissioning team comprised subject matter experts skilled in process technology supported by “generalists” thoroughly conversant in overall plant operations. The merits of a structured team approach to brainstorm and resolve commissioning challenges as opposed to direction from individual subject matter experts is discussed. IntroductionStartup or commissioning is a unique period in the life of a project. Each stage of a project-exploration, feasibility, design, construction, commissioning, initial operations and continuing operations-requires special talents, know-how, and practices. Each stage makes particular demands of the people involved. Initial operations are no exception.Having experienced the 1997 startup phases of oxide and sulfide facilities for a gold extraction plant at a remote island in the Southern Pacific, I can attest to the many challenges and rewards. Commissioning for both facilities was extremely successful and brought the entire project on line well ahead of the targeted completion dates. Using this specific project as an example, this paper will highlight some of the key features of project commissioning. In this paper, the term “pre-commissioning” denotes all activities that lead to mechanical completion; “commissioning” occurs when feed is introduced; and “initial operations” are defined as the duration when operations accepts the system/facility and takes over."

Jan 1, 2000

By Guillermo Turner-Saad

"Best practice in the economic evaluation of mining projects and operations is based on the integration of geological, mining, metallurgical, environmental, marketing, economic and corporate information and strategies. Fundamentally, the economic evaluation takes into account mining, blending, stockpiling, processing, smelting, refining and marketing ore resources with grades estimated by geostatistical methodologies.Nevertheless, a realistic economic evaluation consists of a concurrent and integrated optimization process that maximizes the concentrates or products of liberated and selected ore minerals from mineral deposits. The optimum solution is obtained through a trade-off among mining, blending, stockpiling and processing ore reserves with distinct mineralogy and texture characteristics. The concurrent optimization determines both the practical mining sequence of ore reserves and the dynamic and robust mineral processing flowsheet design over the expected life of mining projects and operations.A combined risk-adverse and concurrent optimization process takes into account the spatial uncertainty of the mineralogy and texture characteristics of ore reserves. Additionally, a variety of mineral processing linear, non-linear, additive and non-additive parameters modeled spatially. The spatial uncertainty of these ore reserves characteristics and parameters has a direct impact in the expected mining sequence, mineral processing flowsheet design, concentrates or products quantity and quality, and economic value of mining projects and operations.Geometallurgical domain constrained stochastic geostatistical methodologies and multivariate mathematical models are applied for spatial uncertainty modeling of ore reserves characteristics and parameters. These mathematical models correlate the mineral processing liberation and selectivity responses of the mineralogy and textures within each geometallurgical domain. These models, in turn, are defined based on results of mineral processing and environmental testwork of unbiased and representative samples chosen for their variability. These samples are selected within mutually exclusive spatial geometallurgical uncertainty domains with similar mineralogical and textural characteristics in mineral deposits.Geometallurgical domains are defined by applying and combining multivariate statistical analysis and implicit modeling methodologies. These involve determining and constraining mineral processing liberation and selectivity properties explicitly contained in the extensive and integrated exploration drillhole database."

Jan 1, 2010

By G. Warren

"In the past, smelting and refining of the precious metals recovered from ore has been generally looked upon as a small and relatively minor component of the business of gold mining. This is not surprising when one considers that the manpower, space and capital components of a gold refinery at an operating mine can be several orders of magnitude smaller than other unit operations at the same location. This philosophy is changing however, as economic and environmental pressure are forcing a re-evaluation of all aspects of gold mining and milling.This paper describes and evaluates three furnace technologies available and used by precious metal operators. The operation of arc, induction, and reverberatory furnaces are discussed and their strengths and weaknesses with respect to capital cost, impurity removal, slag quality, and environmental concerns highlighted.INTRODUCTIONIn the past, smelting and refining of precious metals recovered from ore has been generally looked upon as a small and relatively minor component of the business of gold mining. This is not surprising when one considers that the manpower, space, and capital components of a gold refinery at an operating mine can be several orders of magnitude smaller than other unit operations at the same location. Generally, gold producers will spend a great deal of effort to squeeze the last drop of efficiency out of a leach or grinding circuit via simulation or testwork but will design and build the refinery based upon past practices or industry standards. That past practices or industry standards will provide a functional refinery is undisputed; that this philosophy achieves the efficiencies that are possible or demanded in other plant operations is debatable.Environmental and worker hygiene requirements are considerations as well. Pollution control equipment can become a substantial contributor to both the capital and operating costs of a refinery and furnace selection has a marked influence upon the type and cost of this equipment. Slag characteristics and subsequent processing options available also play a role in furnace selection as precious metal slag losses and the recovery effort can greatly influence the cost / benefit analysis of a particular technology."

Jan 1, 2001

By Bill Shand

"Placer Dome - Campbell Mine is a gold mine located at Balmertown in the Red Lake District of North Western Ontario. During the last 50 years, the mine has produced over ten million ounces of gold.A change in ball mill grinding media was made in August 1998 from 1.5” diameter forged balls to 1.5” diameter, 18% chrome balls. This was done to reduce grinding media consumption and to potentially help flotation performance.A marked ball test done in September-October 1999 showed that, of the alloys tested, 18% chrome was the optimum for the ore at Campbell Mine. The test showed that consumption of chrome balls should be 3 times less than forged balls.In December 1999, the ball size was increased from 1.5” to 2” in an effort to reduce the amount of oversize rock discharging from the ball mill. This change in size was successful.Grinding media consumption for the ball mill decreased 73% from 1.20 kg/tonne before the use of chrome balls to 0.327 kg/tonne after the use of chrome balls. This clearly indicates that, for the ore at Campbell Mine, there is significant savings in grinding media by using chrome.Grinding ball costs were reduced by 59% from $O.88/tonne of ore milled to $0.36/tonne of ore milled representing a saving of $298,000 per year. This also clearly indicates that using chrome balls is a benefit at Campbell Mine.Test work to show that chrome grinding media is beneficial to flotation has been inconclusive so far but is still ongoing.No tests have been done to see if there would be a benefit with an even higher chrome percentage in the grinding media. Test work in this area is also planned for the future"

Jan 1, 2001

By K. Honkivaara, S. Junnikkala, J. Aho, K. Föhr, P. Lampinen

"Bubble area flux has been claimed to be a major tool to dimension flotation equipment. The paper describes laboratory scale (batch and 301 continuous cell using plant feed) and small-scale (1 ,8m3) industrial experiments with coarse-grained massive sulphide ore.Samples were taken with different froth thickness and air addition rates while the bubble size was measured with an UTC capillary device. Problems of averaging the bubble size measurements and the bubble area flux measurements are also discussed.Material balances were calculated using the commercial program BILCO. The flotation rate constants were calculated for pulp and for froth assuming that we can extrapolate the flotation rate constant of the pulp phase from tests using very thin froth thicknesses.The results indicate that the bubble area flux is not alone an adequate dimensioning tool. With fine sizes there was a nearly linear relationship between the bubble area flux and rate constant for constant impeller revolutions. For coarse sizes the form of the relationship is non-linear and seems to go through a maximum value. Thus the final flotation outcome is not a simple function of bubble area flux but depends on the particle size distribution and other more traditional flotation cell dimensioning variables. The effect of the froth residence time in the total recovery of coarse particles is large, but will not be treated in this paper."

Jan 1, 2000

By Chris Bennett, Glen Kosick

"Process modeling and equipment sizing procedures for SAG circuits has advanced significantly over the past two decades. However, no matter how exact the hardness tests, models and scale-up calculations, if the ore samples being used for design do not accurately represent the orebody then major design errors can be encountered. This paper will describe a hardness testing method for identiQing the power requirement profile of an orebody, to an extent that design errors due to poor sampling can be greatly reduced or avoided altogether. It will provide some insight into the magnitude of hardness variability that can be encountered in an orebody. Finally, the paper will preview how detailed hardness distribution data can be incorporated into mine block models, to optimize SAG circuit design for an orebody, both fi-om a capital investment and throughput point of view.INTRODUCTIONIt has long been known that many SAG/AG' circuits experience wide swings in throughput. These throughput variations are almost always of a magnitude not anticipated in the plant design; consequently, they are detrimental to the expected return on investment for the mining companies. It is clear that the throughput swings are mainly caused by changes in ore hardness. In 1993, in an attempt to address this problem, MinnovEX began investigating the possibility of using a small scale SAG test (2 kg charge) to predict SAG power requirements from drill core samples. The objectives of the small scale test were to enable mining companies to do many more SAG hardness tests than currently undertaken for a particular orebody. To make this possible, the SAG tests had to use a minimal amount of sample, be cost effective, and give acceptable accuracy in the prediction of SAG power. If this small scale SAG test proved accurate, then the high sampling frequency would yield results that approach the true hardness distribution of an orebody. This hardness distribution could then be used to derive the distribution and variability of specific energy required throughout the complete orebody. This specific energy (kWh/t) can be thought of as the power required to grind one tonne of ore per hour (kW/tph) to a specified size. Once the true distribution of power requirement throughout the whole orebody is understood, the grinding circuit design team can make an informed decision regarding the hardness value that is most appropriate for sizing the SAG equipment. Throughput losses due to hardness variation could be deliberately planned for or avoided altogether (a capital cost decision)."

Jan 1, 1999

By M. Pourkani, M. Karga, Samid Banis, Andre Laplante, G. Langari-Zadeh

"The Sarcheshmeh mine processes 40,000 t/d of ore grading 0.9% Cu and 0.035% Mo. Grinding is effected in single stage ball mills using 80 mm forged steel balls, with an average consumption of 850 g/t. One of its 5 m diameter, 8 m length ball mills was emptied to measure the ball size distribution. The broken balls were first separated from the total charge, and the remaining balls classified into five size intervals of +7O mm, -70 +60 mm, -60+50 mm, -50+40 mm, -40 mm.The charge was composed of 280,534 balls (excluding broken balls) with a total mass of 305,180 kg. It was found that 4.0 % of the balls were broken; the size distribution of the charge was also markedly different from the desired one, in particular for the small-sized balls. For instance, 9% of the balls were expected to measure 2.54 to 3.18 cm in diameter, but none was found. On the other hand, the amount of top size balls was 30% higher than expected. The radial hardness of balls, which was determined by cutting the balls in halves, showed that the hardness sharply drops when the balls become smaller than 65 mm. The unexpected ball size distribution was attributed to the rapid change in the balls radial hardness. Based on the measured ball size distribution, a new ball make-up charge which uses 80% of 80 mm and 20% of 60 mm balls instead of 100% of 80 mm was proposed and implemented. The effect of the new ball make-up procedure in terms of ball consumption and grinding efficiency is being investigated."

Jan 1, 1999

"The McArthur River uranium deposit is the single richest uranium ore body developed in the world to date. The ore body’s average grade is 15% U3O8. Proven reserves are in excess of 450 million pounds. The grinding circuit is located underground. The ground ore is pumped to a surface facility where the slurry is thickened and loaded into containers for shipping to the Key Lake mill 78 kilometers to the south.The paper will briefly describe the mining method employed. The underground ore handling and grinding circuit, and surface facilities including slurry load out, wastewater treatment and the freeze plant will be described in detail. The paper will emphasize radiation and safety procedures that are followed to minimize operator exposure to radiation risks.LOCATIONThe McArthur River mine is located 78 kilometres north of the Key Lake mill. It is located at latitude 57’ N in northern Saskatchewan, within the southern boundary of the Athabasca formation. The nearest town is Pinehouse, located 320 kilometres to the south. Access to the mine is by an all weather road and by a private airstrip.HISTORYThe McArthur River Operation is jointly owned by Cameco Corporation (69.805%) and Cogema Resources Inc. (30.195%). Exploration for uranium in the McArthur River area started in 1980 by the Saskatchewan Mining Development Corporation, a predecessor company of Cameco.In 1984, a 12-kilometre long basement conductor was identified at a depth exceeding 500 metres. Systematic exploration drilling of this conductor resulted in the discovery of the P2 North deposit. Subsequent surface drilling programs from 1989 to 1992 delineated a narrow ore zone over a strike length of 1700 meters at a depth of 500 to 600 metres. At this time, geological reserves for the deposit were estimated to be 2.4 million tonnes at a grade of 5.0% U3O8 for a uranium content of 260 million pounds U308"

Jan 1, 2001

By Antônio Peres, Rogério Guimaraes, Lauro Takata, Harold Wyslouzil

"Serrana’s concentrator, in Araxii, has capacity of producing approximately 800,000 tomes of phosphate rock per year, with a P205 grade of 36%, from ores grading between 13% and 14% P205. The annual mining rate is around 3.5 Mt and the apatite is concentrated according to the following sequence: crushing, grinding, desliming, flotation and low and high intensity magnetic separations.Two industrial flotation columns (3.0 m x 4.5 m x 14.5 m) were commissioned in 1993 as a replacement for circuits of barite and apatite fines flotation. In 1994, four additional flotation columns of the same dimensions were put into operation, replacing 42 Wemco mechanical cells (300 ft3 each), in the barite pre-flotation, coarse apatite (two columns) and tailings regrind flotation Circuits. Overall, 66 mechanical cells (300 ft3 each) were replaced by only six flotation columns, designed by Cominco Engineering Services Ltd - CESL.The present paper describes the stages of process development, pilot plant testwork and implementation of the industrial project, highlighting the most relevant aspects of each of them. The results achieved are related to improvements in flotation selectivity, gains in P205 recovery and better process control, resulting in better quality products and lower production costs. Other benefits included a decrease in flotation reagents, electrical energy and maintenance materials consumption. These factors contributed to a significant decrease in the cost of phosphate rock production at Serrana, in Araxa - M G."

Jan 1, 1999

By Paul J. Sylvester, Yeonuk Choi, Mike Nelson, Louis J. Cabri, Mike Tubrett

"The development and characterization of sulphide standards containing as many as seven precious metals has increased the efficiency and comprehensiveness of micro-analyses for deportment studies using laser ablation microprobe inductively coupled plasma mass spectrometry (LAM-ICPMS). Because trace precious metals occurring in solid solution or as sub-micron inclusions in sulphides, arsenides, and sulpharsenides are often heterogeneously distributed, a significantly large number of analyses are required. Using well-prepared samples it is possible for up to 120 analyses per day, far more than by other comparable methods. Examples will be discussed from projects focused on the deportment of Au and the platinum group elements (PGE).INTRODUCTIONBecause gold and the platinum-group elements (PGE) occur as both discrete minerals and in trace quantities in a variety of minerals (mostly sulphide, arsenide, and sulpharsenide minerals), specialized trace analytical methods were developed and applied to determine the deportment of precious metals in ore deposits, requiring sensitivities ranging originally from about hundreds to presently tens of ppb for individual precious metal or less.In mineralogical studies, micro-analytical techniques such as the electron microprobe were first applied to obtain analyses for major elements in about 1959-60, but as the instrumentation improved early attempts for PGE in potential carrier minerals indicated that reasonable levels of detection (LOD) were of the order of 300-400 ppm, depending on analytical conditions, X-ray spectral lines, and mineral matrix (e.g., Cabri and Laflamme, 1976). Eventually, the lowest detection levels by electron probe micro-analyses (EPMA) obtained for some PGE were of the order of 10s of ppm using high accelerating voltage and current for extended measuring time on peak and background (Gervilla, Cabri, Kojonen, Oberthür, Weiser, Johanson, Sie, Campbell, Teesdale and Laflamme, 2004)."

Jan 1, 2010

By Donald Leroux, Lori Lynn Smith, Robert Hutchison

"The grinding circuit of the Heath Steele concentrator was sampled under five different sets of operating conditions. The balanced results of the sampling campaigns were used to calibrate steady-state rod mill, ball mill and cyclone models using a commercial simulation software package (JKSimMet® ). Thirteen process variables had the potential to affect significantly the performance of the grinding circuit. Statistical-Design-of- Experiments techniques were applied to plan the suite of simulations in order to improve the quality of and minimize the time to find the optimal solution. The effect of each factor on three performance indices (the cyclone overflow size distribution, the circulating load and the cyclone underflow density) was modeled with empirical polynomial relationships using another software package (Design-Expert@)T. he conditions under which the cyclone overflow size distribution was maximized while respecting circulating load and cyclone density constraints were identified using these relationships. Among the most interesting results is the prediction of a set of conditions where the circuit target size distribution is achieved with only three ball mills instead of four.IntroductionProduction at the Heath Steele concentrator (near Miramichi, New Brunswick) started some 43 years ago at a rate of 750 t/d. The plant capacity was increased to 4000 t/d through successive expansion phases in the seventies (Sapiro, 1977; Johnston, 1978). However, as a result of depleting ore reserves, the plant throughput had to be gradually reduced to the current production rate of 2700 t/d. Heath Steele Mines currently have proven and probable reserves which would see it producing until the end of 1999 (Noranda Inc., 1997).An analysis of the grinding circuit of the Heath Steele concentrator was recently commissioned to identify opportunities for reducing operating costs while maintaining budgeted mill throughput and metallurgical performance. Since the plant will likely be operated below nominal capacity for the rest of its lifetime, the possibility of improving grinding efficiency by operating with fewer ball mills and/or crushers was contemplated."

Jan 1, 1999

By Mike Ounpuu

"Locked cycle testing has become more common practice in the last 10 years. Reasons vary from a desire to avoid pilot plant testing, demand for more variability testing within deposits, and perhaps most importantly, a request from the financial institutions who invariably finance many of the projects. Surprisingly, many people don’t understand locked cycle testing, analysis and balancing. This paper provides a brief overview of locked cycle testing, and then progresses into a more detailed discussion on the analysis and balancing of locked cycle test data. The paper endeavors to educate CMP personnel on how to determine if the test was good and if a reasonable metallurgical projection was made. The paper uses case studies to illustrate the key points.INTRODUCTIONA locked cycle test is a repetitive batch used to simulate a continuous circuit. The basic procedure has a complete batch test performed in the first cycle which is then followed by similar batch tests which have “intermediate” material from the previous cycle added to the appropriate location in the current cycle. These batch tests, or cycles, are continued in this iterative manner for an arbitrary number of cycles. The final products from each cycle, i.e. final concentrate and final tailings, are filtered and thus removed from further processing. At the end of the test, all the products, final and intermediate, are dried, weighed and subjected to chemical analysis. The test is balanced and a metallurgical projection is made.This description can be found in many of our classical textbooks for the mineral engineer 1, 2, 3,4. Unfortunately, none of these textbooks provide any discussion beyond the basic procedure. Often, they provide comments such as in Taggart’ where he states; “It is questionable whether in any case it approximates mill results any more closely than the standard batch test”. It is truly surprising that our classic textbooks promote that locked cycle tests are more art than science, and suggest that they can be of dubious value. None of the textbooks provide meaningful insight or discussion in how to assess:"

Jan 1, 2001

By Andy Cole, Ken Thomas, Stephen Durn, Jacques McMullen

"The roasting and gas handling circuits of the Barrick Gold’s new Roasting Facility at the Goldstrike Mine in Nevada, USA incorporate well proven technologies in a unique combination. During the design of the plant, every effort was made to utilize the best available technologies from the perspective of capital, environment and operating costs.Product from the up-stream dry grinding circuit requires roasting to maximize gold recovery. The roaster utilizes Freeport-McMoRan whole ore roasting technology with pure oxygen as the oxidizing gas. This process was chosen based on its proven application on Carlin-type “double refractory” ores. Double refractory is indicative of the gold being difficult to direct leach due to a combination of pyrite encapsulation and the carbonaceous preg-robbing of the ore.A very extensive gas cleaning train follows the roaster and comprises direct cooling, dust scrubbing, indirect cooling, wet electrostatic precipitation, mercury scrubbing, SO2 scrubbing, CO incineration and NO, reduction before venting to the atmosphere.The roasted calcine is quenched, neutralized and then thickened before leaching in the carbon-in-leach-circuit.INTRODUCTIONBarrick’s strategic development plan integrated the production of gold from Goldstrike’s carbonaceous ore reserves by 2000. A due diligence review of carbonaceous ore processing options was initiated during the second half of 1 996. Processes considered included autoclaving followed by a proprietary ammonium thiosulfate leaching process, fine grinding/autoclaving as practiced then by Santa Fe, and whole ore roasting as practiced by then Independence Mining and Newmont. The objective by year-end was to recommend the most robust and economical process for Goldstrike’s carbonaceous resources. In the fall of 1996, a detailed review of roasting as a processing alternative for Goldstrike was initiated."

Jan 1, 2001

By Michael Hood

An overview is given of some of the major mining research activities in Australia. Automation of mining equipment is not an objective in its own right. The objective is to improve the safety and the economic performance of mining operations. Automation can be a useful method for achieving this objective. There is a strong focus on mining equipment automation and related issues and specific examples of ongoing and recently completed automation projects are described.

May 1, 2002

By Arnaud Faucher, Claude Bazin, Marilène Renaud, Mohamed Mai Manga

"Grinding circuits are operated in the presence of disturbances due to ore changes and equipment malfunctions. One of the challenges faced by the operators is the detection of a disturbance, the identification of its origin and of the action needed to minimize its impact on the process. Hydrocyclones overloading, worn or partly plugged apexes of the units of a nest, surging and internal liner detachment may cause a loss of classification capacity and quality. Recovery losses in downstream leaching or flotation circuits justify the effort in developing techniques for the detection of hydrocyclone malfunction. Two approaches for the detection of hydrocyclone fault are presented in this paper. The first proposed method uses the conventional sensors available in a grinding circuit allowing real time detection of hydrocyclone abnormal operation. The second technique is based on the use of vibration sensors installed on the hydrocyclone of a nest. This last technique enables the classification of the hydrocyclone malfunction as surging, overloading and plugged apex. The two techniques can be combined to increase the robustness of the problem detection.INTRODUCTIONHydrocyclones are used for many purposes ranging from water separation, fines removal and obviously for the preparation of the feed for subsequent concentration circuits in a mineral processing plant. The strategic role of the hydrocyclones for the preparation of the feed to the concentration circuit is often underestimated because of their simplicity of operation and robustness (Napier-Munn et al., 1996; McIvor, 1988). It remains that despite their simplicity, hydrocyclones have been studied extensively from 1970 to 1990 with the objective of developing mathematical models of their operation to be incorporated into grinding circuit simulators (Napier-Munn et al., 1996; Lynch, 1977; Flintoff et al., 1987). Such a development effort is already an acknowledgement of the importance of these units in a mineral processing plant. Several authors (Del Villar et al., 1996; Plitt and Kawatra, 1979; Casali et al., 1998) proposed soft sensors to estimate the classification efficiency of hydrocyclones using sensor measurements. However, few of the proposed sensors are used in the industry. More recently several papers were published to discuss the application of 3-D modelling of the hydrocyclones (Narasimha et al., 2005; Cullivan et al., 2004) and sensors (Bond et al., 1999; Neesse et al., 2003; Hou et al., 1998; Olson and Waterman, 2004; Williams et al., 1995; Willams et al., 1997) to help in the detection of hydrocyclone malfunction."

Jan 1, 2010

By Michael Young, Josh Rubenstein, Michael Cooper, Tom Shouldice, Dan Curry

"Xstrata Nickel Australasia’s Cosmos plant in Western Australia is currently implementing a debottlenecking project to increase nickel output while maintaining metallurgical performance and a low unit cost for production. Mineralogical analyses indicated that the processing bottlenecks could be removed by simply reconfiguring portions of the circuit as well as adding capacity with technical advancements – notably an IsaMill™, Jameson Cell, Wemco® SmartCell™ and Larox Filter.The Wemco® SmartCell™, which is particularly suited for coarse mineral flotation, and Jameson Cell, which is particularly well suited for fast floating and fine minerals, combine as rougher scalpers to complete the main stream portion of the flotation circuit. Addition of the two new cells meet the required flotation capacity to handle the forecast increase in nickel feed grade and tonnage.The addition of an IsaMill™ for regrinding allows the SAG mill to produce a coarser primary product, thus enabling increased throughput. The advantages offered by the IsaMill™ include operation in open circuit (internal classification), an inert grinding environment (ceramic media) for better downstream flotation (“clean” new surfaces), compact footprint (easy to retrofit into existing plant) and energy efficiency. With the new rougher capacity addition, the remaining original plant cells were reconfigured to suit the new regrind mill and cleaning requirements. The existing 24 m2 Larox pressure filter was replaced with a larger 32 m2 unit to dewater the increased concentrate output.The paper describes the upgrade at the Cosmos concentrator with particular emphasis on the mineralogical data and metallurgical benefits associated with the changes.INTRODUCTIONThe Cosmos concentrator is located 680 km northeast of Perth in Western Australia. The deposit’s development and operations were started in 1999 by Jubilee Mines NL, Cosmos’ previous owner, with the first production of nickel concentrate in April 2000. The Cosmos concentrator treats a komatiite hosted high grade nickel sulphide ore, with sulphides being predominantly pentlandite and pyrrhotite with lesser amounts of pyrite and chalcopyrite."

Jan 1, 2010