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By Martin Plante, Denis Hamel, Marie-Pier Bédard, Jeff McLaughlin

"The results and methodology for the optimisation of the zinc flotation circuit at the Matagami Mill using the 6-Sigma methodology are shown in detail while only the base line and final results are presented for the copper circuit. As a result of the six sigma project, the zinc concentrate grade was improved by 2% and the zinc recovery by 3%, with reduced dosage of reagents while the copper concentrate was improved by 2.4%. The process variation in the zinc concentrate grade and recovery was also reduced by over 50%. The main factors contributing to the project success were identified to be improvement in the plant measurement systems, an internal benchmarking exercise and statistically designed laboratory and in plant tests to guide the process improvement. The annual value to Noranda resulting from this project has been estimated at approximately $5 million.INTRODUCTIONThe Matagami concentrator started ore processing from the new Bell- Allard deposit in July 1999. Concentrator performance (grade and recovery of zinc and copper concentrates) was highly variable and work was ongoing to fix the sources of variation. In the mean time, Noranda embarked on its 6-Sigma initiative across the company. The 6-Sigma DMAIC (Define, Measure, Analyse, Improve, Control) methodology was used in the two flotation circuits to reduce variation and improve performance. The statistical 6-Sigma methodology moves towards identifying and controlling critical process operating variables (the X' s) to achieve customerspecified performance (the Y's). The goal of the methodology is to describe a process in a numerical equation known as the transfer function, given as Y = f (XI, X2, X3.. .), where Y could be concentrate grade or recovery and examples of X's are reagent dosage, particle size, etc. With the transfer function, optimal operating conditions and control tolerances are established for the key X's to ensure the desired Y performance is obtained.Each phase of the 6-Sigma methodology is explained below with examples from the Matagami project."

Jan 1, 2001

By C. H. Lee, K. Adham, D. Lemieux, J. P. Mai, A. M. S. Hussein, S. Francey

Copper concentrates contain an increasing amount of arsenic, as the older sources of cleaner ores are depleted. At the same time, the market-determined price penalties on arsenic rich concentrates are rising due to limited arsenic treatment capacity at smelters and increasingly stringent environmental emissions limits, necessitating arsenic removal from the copper concentrates. Heating the concentrate to about 700°C can remove significant amounts of arsenic (as elemental and/or sulfide); however, it is important to achieve that removal economically and with due consideration to post-roasting arsenic (toxic waste) treatment. A comparison is made among different roasting options (kilns, multi-hearths and fluid beds) to qualify their potential application to arsenic roasting of copper concentrates. Both direct and indirect methods of heating as well as inert, air and oxygen atmosphere options are considered. Technology comparisons are made with respect to the roaster’s operation, its off-gas treatment and arsenic stabilization. INTRODUCTION In the last twenty years, the arsenic content in globally produced copper concentrates have risen steadily due to the depletion of cleaner ores and the increase in complex concentrate production. Smelter contracts typically limit arsenic to less than 0.5wt%, with a penalty paid per unit of additional arsenic and other contaminants such as antimony. While penalties can vary, they can be on the order of $70/dry metric tonne of concentrate per 1% of arsenic. Strategies used by miners to avoid the price penalties have been to blend clean and complex concentrates, and to remove arsenic from complex concentrates to achieve smelter limits. Arsenic can be removed from copper concentrates by the heat-induced decomposition of the predominant enargite mineral (Cu3AsS4) at ~700°C. Roasting at such elevated temperature is a commercially proven option, but still suffers from process, cost and environmental challenges. After roasting, the released arsenic sulfide fumes can be separated from most of the entrained concentrate dust in cyclones. Then, the As2S3 laden off-gas can be oxidized to generate a still volatile oxide (As2O3), which can be further separated from the residual concentrate dust in a hot electrostatic separator. Next, the off-gas can be cooled with air or water, to either deposit the arsenic oxide in a dry or dissolved state. Dissolved arsenic acid can be oxidized to As(V) and reacted with iron to yield a stabilized ferrihydrite precipitate or scorodite that can be safely disposed. In the case of dry arsenic trioxide, it can be mixed with reactants (e.g. iron) and fluxes before thermal stabilization in a glass matrix, thus eliminating the need for costly oxidation.

Jan 1, 2019

By H. D. B. Wilson, W C. Brisbin

The Thompson-Moak Lake nickel belt lies along the root of a Precambrian mountain range of the island-arc or alpine type. This range coin-. cides with the boundary between the Superior and Churchill blocks of the Canadian Precambrian shield. The nickeliferous peridotites and nickel ores occur along a negative gravity strip, which also coincides with strong thrust faulting and the axis of greatest deformation of the mountain range. Similar structures and highly nickeliferous peridotites occur in present-day island arcs. Much of the belt is covered by Paleozoic rocks, but a length of 270 miles may be explored by present-day methods. Nickel deposits have been located along 60 miles of the belt. The length of the favorable structure and the size of the known deposits suggest that this will be one of the world's major nickel regions. The regional gravity anomalies, which consist of a branching negative strip bordered by two positive strips, are best explained by major warps of a crust consisting of a granitic and basaltic layer. The ore deposit

Jan 1, 1961

By J. E. Terry

This paper outlines the steps that are being taken to close the gap between training and industry and indicates areas in which more work needs to be done. Although based on experience in the Maritime Provinces, it is applicable to industrial and educational studies across the length and breadth of Canada.

Jan 1, 1970

By Keith R. Vining

Solid fuel, usually coke or anthracite, is commonly added to ground iron ore fines in the pelletising process to reduce the amount of burner fuel required and the maximum firing temperature, while maintaining pellet strength and the productivity of the pelletising plant. However, the use of coke and coal contribute to the production of greenhouse gases (eg carbon dioxide) and are a source of NOx and SOx emissions. Utilising wood biomass / wood char is one methodology CSIRO Minerals is investigating as an alternative to coke and coal, as it is a renewable energy source that is very reactive, has a low ash content and the potential to reduce emissions of greenhouse gases from the iron and steel-making process. Previous work has established that charcoal can effectively replace coke in the sintering process (Lovel et al, 2005), and the following experimental work aims to demonstrate that it can also provide an adequate replacement for coke and coal in the pelletising process. The paper presents the results of an experimental testwork program comparing the behaviour of fired pellets containing traditional solid fuels and charcoal. The physical properties of the pellets are compared in terms of their compressive strength, tumble and abrasion indices and the reduction properties. A qualitative assessment of the pellet microstructures is also presented.

May 1, 2007

By R. J. MacLeod

"INTRODUCTIONThe removal of water or liquid from slurries and solutions is a procedure generally present in many process industries. This has prompted the development of many mechanical and thermal devices with which, I have no doubt you are al1 familiar.In the mining industry, beneficiation of ores necessitates ' the removal of water or other 1iquid from the concentrate material. The most common method of accomplishing this is by means of thickeners, vacuum filters and rotary dryers, generally all three and in that order;The use of spray drying techniques to obtain the same end has many advantages and these will be outlined here.THE PRINCIPLE OF SPRAY DRYINGA Spray Dryer consists of a cyclone shaped chamber where hot air and atomized concentrate are introduced in the upper section. (Sec figure No. 1). Efficient mixing of these two streams produce dry concentrate and moist spent air, both of which are removed from the base of the chamber.Thermal dewatering involves the product ion of heat and spray drying has no special requirements in this area. Generally a direct fired air heater is used and the products of combustion whether they stem from the use of gas, oil or coal fuel go into the spray dryer along with quench or dilution air which is used to reduce the combustion gases to the desired temperature. This generally lies in the range of 1,000 to 1,600° F. In spite of these relatively high inlet temperatures, the spray drying process eliminates the need for refractory lining or indeed any special heat resistant metals except on the air inlet ducting. Waste heat may also be used even if a significant amount of contamination is present."

Jan 1, 1972

By Steve Custer, Tim Kahl, Sue Parker

"The primary objective of the Barrick Goldstrike expert system is to achieve the desired liberation size for the autoclave circuit while maximizing throughput. In addition, when downstream limitations (autoclave circuit or others) restrict the milling circuit tonnage the expert system further improves grinding product size. Since its implementation, the expert system has provided expected throughput improvements while improving or maintaining circuit performance.Barrick Goldstrike and MinnovEX Technologies Inc. teamed up to install an expert system to control Line I and I1 of the autoclave grinding circuits. The dissimilar equipments of the two grinding circuits coupled to different production objectives, skepticism among operators, and the variability of the ore types processed in each line challenged the implementation.To ensure implementation success, Barrick Goldstrike dedicated four administrators full time to the project. The system administrators included two metallurgists, a process control engineer and a mill trainer. As usual, the control room operators were the experts who provided the operational logic for the foundation of the control system. The administrators and all eleven of the control room operators signed off on acceptance of the system. Current online utilization is 95%.INTRODUCTIONThe grinding circuit for Barrick Goldstrike’s autoclave process consists of two SAG Mill Ball Mill lines. Line 1 grinding circuit is comprised of a 22 by 8 foot, 2500 horsepower Allis Chalmers (AC) variable speed SAG mill, followed by a 12.5 by 14 foot, 1250 horsepower AC ball mill, and a 13.5 by 18 foot, 1800 horsepower Dominion ball mill in parallel. SAG mill recycle from the primary discharge screen (+5/16 inch) is processed through a 1352 Nordberg Omnicone crusher (4.5 foot). A secondary SAG discharge screen (+1/8 inch) closes the SAG circuit. The ball mill circuit is closed with 6 Krebs super D20 cyclones. Line 2 grinding circuit is comprised of a 24 by 12 foot 4000 horsepower Fuller variable speed SAG mill followed by a 16.5 by 30.5 foot 5000 horsepower Fuller ball mill. SAG mill recycle from the primary discharge screen (+5/16 inch) is processed through a 1560 Nordberg Omnicone crusher (5 foot). A secondary SAG discharge screen (+1/8 inch) closes the SAG circuit. The line 2 ball mill circuit is closed with 12 Krebs super D20 cyclones. The underground Meikle ore is typically fed through Line 1. Achieving the desired product size is critical for successful processing of Meikle ore. Open pit Betze ore is fed to Line 1 when Meikle ore is unavailable. Line 2 processes Betze ore exclusively. Throughput is maximized when processing Betze ore as the product size is less critical. The products of these two circuits are combined as feed to the autoclave circuit. The ore types and the autoclave circuit demands determine the required grinding circuit product size and maximum combined mill feed tonnage."

Jan 1, 2001

By U. Olsson, M. Ramström

"The IREDES1 Special Interest Group (SIG) for drill rigs has been active since its launch during MineExpo 2000 in Las Vegas, Nevada, United States. SIG for drill rigs is responsible for the development of an IREDES drill rig profile, including the exchange of production planning, performance, and quality reporting data between various kinds of drill rigs and the databases used by the operating mine. During 2001, a first draft of the drill rigs profile was jointly prepared by drill rig suppliers, mining companies, and third-party suppliers and reviewed during several SIG meetings. The profile for prototyping projects was released in mid-2002; the first official IREDES release was based on feedback received from these projects.In 2002, LKAB and Atlas Copco decided to carry out an IREDES prototyping project using the Simba L6C production drill rig. The goal was to verify the drill rig profile and identify potential problems prior to the official release. This was done by implementing the IREDES prototype version into LKAB’s production planning and Simba rig control systems. The Simba rig control system was designed to handle drill plan and logging information coming from the mine planning system. In parallel to this, the mine production system was adapted for drill plan export and logging data import in IREDES format.This paper presents the most recent results of the IREDES prototyping project. Results from late 2002 show that the basic architecture of the drill rig profile works and that a common language for data exchange, IREDES, will be a central requirement for future mine automation.As a result of the IREDES prototyping project, the IREDES standard was approved at the IREDES board meeting held at the 2003 CIM Annual Conference and Exhibition. Soon thereafter, IREDES compliant products, including drill rigs and mine production planning systems, were put on the market."

Jan 1, 2005

By Patrick Pelletier, Pierre-Alexandre Bossé, Julie Fournier, Gianni Bartolacci, Jayson Tessier, Ahmed Bouajila, Carl Duchesne

"This paper summarizes research into the study of the control of flotation processes based on froth images. The use of numerical image analysis techniques for the extraction of significant patterns (froth structure and colour) from industrial flotation froths is discussed.First, the froth was analysed by the use of the Multivariate Image Analysis (MIA) method to extract the spectral variation contained in a RBG image. The Partial Least Squares (PLS) was applied to develop the empirical model for the prediction of froth grade. The results of this application are illustrated using one industrial case study.Second, the froth was also analysed by Gray Level Co-occurrence Matrix (GLCM) and Wavelet Transform Analysis (WTA) to extract features related more to morphological aspects. Since there is useful information hidden in froth texture, these methods are used as texture descriptor to classify different types of froths.Third, a control strategy based on a froth structure indicator was developed and implemented on an industrial zinc flotation circuit. The feedback controller uses a reagent flow-rate in order to maintain the froth structure at its set point. Process improvements obtained with the implemented control strategy are highlighted.Finally, the feasibility to implement a hybrid control strategy consisting in a froth structure supervision controller based on the froth concentrate grade estimated by MIA and recovery inferred by image analysis is discussed.INTRODUCTIONThe development of control strategies to achieve satisfactory control performance has always been a great challenge for engineers and the need for improved process diagnosis, control and optimization has become more and more evident for the mineral processing industry in the last decades. That is why new process variables need to be measured particularly in highly multivariable processes such as flotation. This industrial process is subject to a wide variety of disturbances mainly generated by frequent and sometimes abrupt changes in physical and/or chemical ore characteristics and also, due to some irregularities in manual process operating procedures. A flotation circuit scheme is always a complex network of more than two dozens slurry streams. Information about those streams (% solid, metal content analysis and flow rates) is important for process diagnosis performance evaluation and control. Also, because of the grade decrease and increasing complexity of the new processed ores, more instrumentation outputs and more powerful and significant pr"

Jan 1, 2005

By André Laplante, Robert C. Dunne

"The role of flash flotation in precious metal recovery is reviewed. Interacting flash flotation and gravity recovery is discussed. The GRG test can be used, when linked to the behaviour of gold in grinding and classification units, to predict the recovery of GRG by flash flotation. The Cadia mill in New South Wales is used as a case study of circuit design, performance and simulation. It is shown that high throughput coarse grind applications such as Cadia require very effective fine GRG recovery systems in the primary grinding loop, which are more consistent with flash flotation than batch centrifuge concentrators (BCCs). BCCs can effectively recover a significant fraction of the GRG from flash concentrates, as predicted by laboratory testing and confirmed at a growing number of mill sites.INTRODUCTIONThe GRG test was designed to characterize Gravity Recoverable Gold (GRG) ten years ago (Woodcock and Laplante, 1993; Laplante et al, 2000), in response to the need to predict how much gold could be recovered from the main circulating load of grinding circuits using centrifuge concentration. The amount of GRG is not itself a prediction of gravity recovery: the performance of a gravity circuit nested in a circulating load also depends on the type and size of unit used, how much of the circulating load is treated, and how many times GRG is recycled and presented to the recovery circuit. The GRG test was first used in a simulator to predict gravity recovery (Laplante et al, 1995), and indeed most of the early tests targeted ores or prospective ores for which gravity recovery was an obvious match. As the test became more accepted, more and more samples returned a verdict that precluded the use of gravity recovery, or made it marginal. Typically, such ores are either copper-gold or base metal ores in which the GRG content is often fine (largely below 105 um) and ranges from 8 to 33%. Extensive simulation and plant sampling tests, either published (Laplante et al, 1997) or proprietary, show that full scale gravity recovery for these applications is generally below lo%, and can be as low as 2%. Not in all cases are such low recoveries economically unjustifiable, since there is documented evidence (Darnton et al, 1992) that a significant proportion of the gold recovered, at least in the case of base metal ores, would not have reported to the copper or zinc concentrate and received NSR credits."

Jan 1, 2002

By Leslie F. Sawatsky

Raw water import requirements for oils sands mines depends mainly on factors such as ore processing rate, ore material characteristics, tailings processing technology, tailings storage systems, process water storage operations, the phase of tailings production, and non-recoverable seepage losses from tailings ponds. These factors can be readily quantified to calculate raw water import needs. Nevertheless, mine water balance continues to be misunderstood by some regulators, stakeholders and possibly even mine operators. Confusion over consumptive uses has led to a focus on recycling which is not necessarily applicable to a closed circuit water management system. Some water conservationists focus attention on raw water intake rules instead of the technical decisions that control consumptive uses. Typical oil sands mine water balances for various types of tailings operations, are analyzed to illustrate the effect of controllable and uncontrollable parameters as well as major options that need to be explored for reducing water demand. Uncontrollable parameters including tailings fines content, oil content of ore, connate water, mine depressurization water inflows and surface water hydrology. Controllable parameters include storage operations, advance water supply ahead of mining operations, type of tailings operation, net tailings pond seepage losses, and tailings impoundment operations. Potential reductions of raw water supply rates may be obtained by providing greater on-site storage to attenuate fluctuation in water demand, advance water accumulation prior to mine start up, early commencement of CT operations, thickened tailings disposal and paste technology. Water conservation measures should be focused on reducing on-site process water storage and water content of tailings material.

May 1, 2004

By Harry Siewert

"The commissioning of the 4th concentrator (C4) at Freeport Indonesia in early 1998 increased total mill throughput from 130k to 230k tonnes per day. The C4 flowsheet consists of a single 38 ft diameter SAG mill, four 24 ft diameter ball mills followed by four banks of 4,500 ft3 WEMCO flotation cells.The design of the C4 rougher flotation circuit was based on the expectation that the copper and gold recoveries achieved with the 4500 ft3 flotation cells would be equivalent to or better than that achieved with the 3,000 ft3 flotation cells in the 3rd concentrator (C3) rougher circuit. This assumption was due to the fact that the C4 flotation feed size would be less than 15% +212 microns compared to the C3 rougher flotation feed size of less than 23% +212 microns. By mid 1998 it had become apparent that the C4 rougher flotation copper and gold recoveries were below those predicted.In order to determine the cause of the lower recoveries, stage by stage rougher flotation surveys were performed in all four concentrators. The results of the surveys showed that the C4 rougher flotation circuit had a lower recovery of +200 micron material than that achieved in the C 1, C2 and C3 concentrators. Through mineralogical analysis, it was determined that a majority of the copper contained in the +200 micron fractions was liberated and recoverable. This paper describes the steps taken to increase coarse particle recovery, overall plant recovery and the results achieved."

Jan 1, 2001

Safe underground mine backfilling with hydraulic fill requires efficient fill drainage and stable fill retaining walls or bulkheads. If any or both the above two requirements are not satisfied, such a backfilling operation creates unsafe conditions and may lead to backfill retaining wall failures. Mine backfill retaining wall failures are always associated with economic losses, fatalities and interruption to mine production. Mining industry wants to either reduce or eliminate these fill retaining wall failures. The fill retaining wall failures could be effectively avoided through the use of fibre reinforced shotcrete fill retaining walls and engineered drainage systems through the retaining walls. This paper presents the experience of the author in developing, implementing, and monitoring several hundred fibre reinforced shotcrete fill retaining walls at a mine in Australia with the highest safety standard. The details on the mine backfill drainage system and fibre reinforced shotcrete fill retaining wall designs are also presented.

May 1, 2007

By H. A. Dakin

"An introduction to the metallurgical requirements of the Flin Flon concentrator reveals the magnitude of ore blending which is required in its operation of three flotation circuits. The resulting variability in flotation heads which has in-creased the separations complexity justified the development of an on-line analysis system.This paper provides a summary of the development and operation of that system. Particular attention is devoted to the problems encountered and where possible their solutions; as the requirements of accurate information and minimum downtime of the sampling system and x-ray analyzer are considered essential. The addition of a recess central computer has provided information in a more easily understood form and as well has allowed the start of automatic reagent control.An indication of the on-line analysis system's value to the concentrators' operation is provided by an examination of the plant in metallurgical data and reagent costs in previous years.The metallurgical plants of Hudson Bay Mining and Smelting lie adjacent to the city of Flin Flon on the Manitoba-Saskatchewan boundary. Plant operations commenced in 1930 treating copper-zinc ore from the local mine. In recent years ore bodies near Flin Flon and in the Snow Lake area, about seventy-five miles east, have been brought into production.By the beginning of 1973 the number of producing mines had increased to nine; with all the ore being handled by the Flin Flon concentrator. Figure l, showing mine ore assays during 1973 indicates the variation in metal content possible."

Jan 1, 1974

By Peter Radziszewski

"Comminition process costs, which typically represent an important contribution to mining operating costs, can themselves be divided roughly between energy and liner /charge wear. As such, developing the ability of predicting mill wear will allow a more accurate estimation of wear costs and therefore their control. In this paper and based on the possibility of estimating energy dissipated in impact and abrasion by using a discrete element charge motion model for mill charge motion simulation, media wear will be investigated principally as a function of three wear mechanisms : impact, abrasion and corrosion. Experimental results will be used to identify a total media wear model. Model results will be compared with real mill data from three operating Australian mines. These preliminary results show for the cases tested that the overall average error, as compared with the observed cases, is 0 % with a standard deviation of 66.1 %. Results from the Bond abrasion index tests, for the same cases investigated, give an overall error of 73 % and a standard deviation of 192.5 %. Further work is in progress to improve test procedure, validity and confidence.INTRODUCTIONComminition process costs, which typically represent an important contribution to mining operating costs, can themselves be divided roughly between energy and line / charge wear. In Greenfield situations where the estimation of operating costs is critical in determining project feasibility, the level of confidence in such estimates becomes quite important. With regards to media wear, the Bond abrasion test currently serves as the standard media wear prediction device. However, as shown [l, 21 test results once compared with actual wear rates can vary greatly. Figure 1 shows test results as compared with 15 cases. Average error over this data is 73 % with a standard deviation of 192.5 %. As such, developing the ability of more accurately predicting mill wear will allow greater confidence in wear cost estimation and therefore their control."

Jan 1, 2001

By David W. H. Ellis

This presentation will discuss over 2 dozen past, present and future oils sands projects in Alberta. The very early projects as well as the most recent along with the ones that never got off the ground and the ones in production are all included. For each, the following will be given (where known): owners, location, type (mining/SAGD/refinery), production, capital budget, spent to date, status. Some overall impacts will be discussed.

May 1, 2004

By K. Leung, J. E. Udd, L. B. Geller, F. Kitzinger

"Sixteen miners died tragically at the Paymaster gold mine in Timmins, Ontario, in February 1945 when a badly corroded hoist rope broke. The investigation of the Royal Commission that followed arrived at a number of recommendations (The Legislative Assembly of Ontario, 1947) aimed at avoiding future disasters of this kind. One consequence was the development and legislated use in Canada, several decades ago, of powerful electromagnetic (EM) instruments for the non-destructive testing of mine shaft wire ropes. The authors have been involved for many years in the design, production, and testing of EM instruments (Geller and Udd, 1990, 1992; Geller and Kitzinger, 1991; Geller et al., 1992). Since the mid-1990s, particular attention was paid to the design and operational results obtained from the computer controlled portable MagnographTM II wire rope tester (Geller et al., 1995, 1998).In the present paper, the authors summarize further hardware- and software-related developments of this instrument. Further, an account is given of the results obtained from the most recent design known as the Perma- ScanTM testing system. This is also a computerized dual-function system. In it, however, the sensor head is designed primarily, if not solely, for permanent installation and for remotely controlled operation. This represents a basic departure from traditional designs, which were intended for ad hoc temporary installations. The remote-controlled instrument described in this paper has now been in operational use for more than four years."

Jan 1, 2003

By Jim Wickens, Vince Lofftus, Mark Kravetz

"This paper discusses how Bald Mountain Gold Mine (Ely, NV) dramatically reduced mercury production by treating process solutions with a liquid reagent based on organic sulfide chemistry. Cherokee Chemical Co. Inc. (Vernon, CA) markets this reagent as UNR 81 1A.A summary of the bench tests comparing various reagents is given. The ensuing plant trial is discussed and successful mercury removal results are presented. Data is presented to show that early removal of mercury in the process reduces the amount of mercury collected at retort. Data is also presented to show that use of this reagent does not reduce gold recovery. Plant trial treatment costs are discussed. Finally, the potential of this technology to substantially reduce closure costs as compared to traditional mercury removal options is discussed.INTRODUCTIONBald Mountain Mine produces gold. Mercury is a costly by-product of gold production because no revenue is generated from mercury. Traditionally, mercury leaches out with gold and follows the gold throughout the processing circuit until it is eventually condensed and collected at Retort. Minimizing or eliminating mercury production minimizes worker health and safety risks, minimizes environmental liability, and reduces costs. In the long term, minimizing mercury production will make mine closure easier and less costly.Bald Mountain Mine dramatically reduced mercury production by treating process solutions with liquid reagents called UNR 811 and UNR 811A. Several reagents were bench tested before proceeding with a plant trial. The reagent complexes the mercury to form an insoluble organic sulfide precipitate. In independent tests, the precipitate passed acid mobilization tests and is considered stable."

Jan 1, 2001

By Dong-kil Lee

Field tests have been conducted to investigate the characteristics of flow and physical chemistry of acid mine drainage(AMD) in the Hwangji-Yuchang settling pond in south Korea. pH and ferrous concentration of AMD flowed in the settling pond is 6.79 and 25.5 mg/l, respectively. As a results, we found that the ferrous oxidation reacted strongly around the inlet of the pond and the distributions of physicochemical properties is depended on flow characteristics and ferrous oxidation rate of AMD.

May 1, 2011

By Ross Byron, Ken Roberts

"During the fall of 2001 a program was initiated to determine the feasibility of improving the metallurgy in the flotation circuit at the Penhorwood Talc Concentrator located near Timmins, Ontario. Bench scale testwork led Plant staff to believe that opportunities existed to improve the flotation performance. Pilot scale testing was then conducted using two vendors and their respective cylindrical type flotation cells to indicate where these opportunities existed.The results of the test program determined that the replacement of the existing flotation circuit (roughers, scavengers and cleaners) with new flotation cells could be economically viable. Justification was due to a reduction in installed horsepower, maintenance costs and a subsequent increase in recovery and throughput while maintaining product quality. Unforeseen benefits included reagent and air reduction, accurate level control as well as improved working conditions due to the removal of the antiquated equipment and ancillaries.This paper will detail the justification, installation and implementation of the new flotation circuit. A special emphasis will explain the hurdles encountered while trying to install and commission the circuit in an operating Concentrator with minimal interruptions to production.INTRODUCTIONThe Penhorwood mine and Concentrator is 28 yrs old and has gone through many ups and downs with different owners, Johns Manville, Steetley Talc and now Luzenac. It has also been expanded twice, with flotation and grinding capacity being added in recent years.All of the concentration is completed at the mine site while the fine grinding takes place in the Timmins micronizing plant.Luzenac’s Philosophy is to be the world leader in Talc production, technology and sales, having about 70% of the current world High Grade talc market."

Jan 1, 2004