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By André Laplante

The mineralogy of the Joe Mann mine differs from that of other deposits of the Chibougamau area that are exploited by Les Mines Camchib Inc.. It is characterized by very fine first generation gold and very coarse second generation electrurn. Successful milling required adjustments of the processing flowsheet, which includes gravity, flotation, and cyanidation. A study of coarse gold liberation in the gravity/grinding circuit is described. Modifications to the flotation circuit and the reconditioning of a small capacity cyanidation circuit are presented. As a result of the circuit modifications, gold recovery increased from 65 to 81% in flotation and cyanidation. Overall gold recovery increased from 77% to 88-89%. These improvements are mostly attibutable to the recovery of -10 urn gold, whose presence in the Joe Mann ore and original circuit tails had been identified by process mineralogy.

Jan 1, 1989

By G. W. Center

This manual is intended to assist federal, state, and local regulatory and planning agencies, the mining industry and consulting engineers in determining the regulatory, environmental, technical and economic constraints influencing the site selection, planning, construction/operation and reclamation phases of a waste disposal facility. The information presented illustrates that factors affecting an individual site may be simple or very complex. This manual discusses methods for making future land use decisions, determining regulatory requirements, evaluating physical, chemical and structural properties of the disposed waste and selecting appropriate waste disposal methods and control techniques. Procedures and guides in the manual will assist personnel in assessing the complexities of waste disposal on a site specific basis. The major types of waste disposal systems are discussed and evaluated in terms of their advantages and disadvantages and applicability to varying site conditions and waste characteristics. Appropriate references following each chapter provide more detailed data on material presented in the text.

Jan 1, 1982

By Brad Bjornson, Geoff Stevens, Mike Ashar, Cliff Britch

"LocationThe facilities are located 30 km north of Fort McMurray, Alberta (or 470 km northeast of Edmonton, Alberta).Brief Description of OperationThe operation includes the mining of oil sands, extraction of bitumen from the oil sands, and upgrading of the hydrocarbon fraction prior to shipment via pipelines to refineries.HistorySuncor was initially called Great Canadian Oil Sands. It was started by Sun Oil of Philadelphia. Construction began in 1965, with the first production coming on September 1967.Details of Present Milling Practice (Plant 3)Plant 3 recovers bitumen from the oil sand using the Clark hot water process. The basic steps in the process are conditioning, primary recovery, secondary recovery and tailings oil recovery.Oil Sand Feed SystemThe oil sand feed from the mine is transferred to the feed bin in extraction by two conveyor systems. Oil sand analyzers on each conveyor system measure the amount of bitumen and fines in the ore. Cones at the bottom of the feed bin distribute the oil sand to five process lines. Feed flow from each cone is regulated by an apron feeder which transfers the oil sand to conveyors that discharge into the conditioning drums. Weightometers on each of these five conveyors measure the rate at which the oil sand is delivered into the drums. The oil sand feed rate to each drum can be adjusted by varying the speed of the apron feeders."

Jan 1, 2000

By M. Fan, A. Hobert, W. van Dyk

Low flotation recovery and selectivity have been long-standing problems in the treatment of most ultra-fine phosphate ores. This paper investigates the improved flotation of two ultra-fine Brazilian phosphate ores from Itafos’ Arraias project. These ores are predominantly composed of quartz, muscovite, apatite and a Ca-Al phosphate. As a result of fine phosphate dissemination and associated grinding requirements, column flotation was selected for the metallurgical test program to address the inherent limitations of fine particle flotation using conventional flotation cells. Accordingly, both benchtop Denver cell and laboratory column flotation tests were conducted to optimize their respective flotation responses. Improved flotation performance was achieved using column flotation as compared to benchtop mechanical flotation. In column flotation, a P2O5 concentrate grade of 30.5-31.0% was yielded at a flotation recovery of 83.7-85.1% for the 38x0µm generated ultrafines. The benefits of Eriez’ column flotation technology realized from this study include a higher degree of selectivity, increased recovery, and reduced collector dosage as compared to bench-scale mechanical flotation.

Mar 2, 2022

To date, improved thickening of clay mineral dispersions presents major technological and environmental challenge to mining and mineral processing industry worldwide. Flocculant-assisted dewatering processes currently employed during thickening are far from being efficient, requiring further development and process optimisation for dramatic improvement. In this paper, recent studies carried out to modify (i) pulp chemistry, (ii) flocculation performance and (iii) particle interactions to significantly improve the dewatering behaviour of flocculated, model and plant smectite and kaolinite clay dispersions are presented. It is shown that pulp pH-mediated hydrolyzable metal ion hydrolysis and specific adsorption in tandem with high molecular weight flocculant and shear may synergistically act to significantly enhance the dewaterability of both clay mineral pulps. Furthermore, polymeric flocculant structure type is demonstrated to show clear links with floc shear sensitivity, and hence, flocculation performance and pulp dewaterability. The implications and relevance of the findings plant clay mineral waste tailings thickening operations are drawn.

Jan 1, 2006

By Graeme Heyes, Warren J. Bruckard, Michael A. Somerville

"A limited series of laboratory batch flotation tests was conducted on a suite of calcium-ferrite-based slags made in laboratory smelting trials. The smelting tests were conducted as part of a wider CSIRO Minerals’ research project aimed at developing a single-stage continuous copper-making process. The objective of the flotation work was to maximise copper recovery and iron rejection from slags made in the process.The slags treated included a reduced low copper slag, an oxidised high copper slag composite, and a very reduced low copper self-pulverising slag. The copper content of the slags varied from 2.9% to 14% Cu. The copper was present as metallic copper, oxide copper and copper-calcium ferrite phases with the phase mix related to smelting conditions. The slags were crushed, ground, wet-screened at 355 or 210 µm to remove coarse metallic particles of copper, and floated at natural pH (about pH 11) using reagents and conditions appropriate for the selective recovery of the copper phases.In rougher/scavenger tests on the reduced low copper and self-pulverising slags copper recoveries above 92% were obtained with copper enrichment ratios varying between 2.3 and 2.9. For the oxidised high copper slag copper recoveries were poorer - only 47% of the copper was recovered with an enrichment ratio of 3.3. The presence of the slow floating copper-ferrite phases in the slags limited further (selective) copper recovery. Ageing of the high copper slag, that is, a significant delay between grinding and flotation, resulted in a deterioration of the copper flotation response.The critical parameter affecting the flotation of copper from direct converting slags was found to be the form of the copper in the slag. Slags containing a higher proportion of the more desirable form of copper for flotation will provide the best opportunity for high copper recoveries to be consistently obtained, allowing improvements in the economics of the overall copper-making process."

Jan 1, 2003

By Manoj Khandelwal, Minghui Liu, Jian Zhou, Kun Du

The waste rock and tailings backfill into the mined-out areas are the most effective method for solving the environmental pollution and surface disasters for nonferrous metals mines. In practice, the success and availability of backfill operations are dependent on the slurry fluidity and the strength properties of cement backfill. The transport of the slurry through the pipeline to the designated backfilling area relies on its eximious flow properties, while the appropriate strength of the filling body ensures the safe operation of the stope. In this paper, the effects of cement and aggregate types on the slurry fluidity and strength characteristics of cemented backfill are studied in detail, which are often ignored in other pieces of literature. Diffusivity is used as an indicator to evaluate the slurry fluidity. Various slurries whose concentrations ranging from 70%, 73%, 75%, 78%, and 80% are made with different aggregate ratios and cement-sand ratios are tested. It has been shown that slurry fluidity is inversely related to its concentration, but 78% is the “stopping point” for the deterioration of fluidity. The addition of rod-milled sand improves or worsens the cemented backfill (CB) strength depending on the amount of rob-milled sand. The uniaxial compression experiment results on 216 CB specimens produced by different combinations of influencing variables showed that CB specimens made from cement with superior mechanical properties have a higher uniaxial compressive strength (σucs). It has been also found that the effect of aggregate ratio on the CB strength is not singular, but works in conjunction with the curing time and the cement-sand ratio. The longer the curing time and the higher the cement content, the higher the CB’s σucs. To avoid the time-consuming and costly problem of obtaining the strength of the CB from indoor experiments, an SVR model capable of predicting the uniaxial compression strength of CB specimens is proposed, which is optimized by genetic algorithm (GA) and particle swarm optimization (PSO) algorithm. The results of the three performance indexes (MAPE, MSE, and R2) show the superior performance of the GA-SVR and PSO-SVR models and the agreement of the predicted results with the experimental results, which indicate that these two models can accurately predict the σucs of CB.

Feb 3, 2022

By J. Kretschmann, A. Müterthies, T. Rudolph, C. Yang, P. Goerke-Mallet

Catastrophic failures of dams and tailing storage facilities (TSF) have triggered the development of integrated ground monitoring concepts for hazard assessment, risk management and public safety. These concepts are based on knowledge and expertise of persons technically responsible for TSF. The aim is to combine this implicit and explicit knowledge with the data of in-situ- and remote sensors. Risk management for TSF must include a comprehensive monitoring program that integrates data from ground based-, air-borneand satellite-based sensors. Researchers at Technical University Georg Agricola (THGA) in Bochum, and EFTAS, Muenster, Germany, are fusing a variety of sensor data to enhance the capability and performance of TSF monitoring to ensure public safety downstream of TSFs. This paper describes the variety of sensors implemented in numerous research projects. It gives an outlook in terms of planned satellite missions with innovative equipment like hyperspectral sensors. The implementation of the data of these sensors in the analysis will increase the understanding of spatio-temporal impacts of mining activities. Future opportunities are explained by a datacube approach.

Mar 2, 2022

By J Hartog, V Beehan, D Beeson

Treating an ore blend from five different deposits at Newgold’s Peak Gold Mine (PGM) has always been a challenge due to the inherent variability between each of the orebodies.Operating the processing plant at PGM with a ‘normal’ level of process control and a good operations and metallurgical team has allowed the processing plant to achieve above the name plate design at 102 t/h and recoveries in the order of 92.8 per cent for gold and 88.4 per cent for copper.With a company value and culture of striving for continuous improvement, PGM implemented the Manta Cube® system, an advanced process control system, for the semi-autogenous grinding (SAG) mill, gold leach circuit and tailings thickener over a 12-month period throughout 2012 and 2013.The overall benefits of implementing the Manta Cube® system has led to an increase in mill throughput of at least two per cent, improved SAG mill liner life, a ten per cent decrease in cyanide consumption and a reduction in flocculant of at least 30 per cent with no detriment to the gold and copper recoveries.The entire project had a payback period of less than two months, and the combined benefits resulted in an overall increase in revenue of at least $5 million per annum. This is quite significant for a processing plant the size of PGM.This paper describes the methodology, overall results and economic benefit of applying the Manta Cube® system on-site at Newgold’s PGM.CITATION:Hartog, J, Beehan, V, Karageorgos, J and Beeson, D, 2014. Optimisation of the Peak Gold Mine’s processing plant through advanced process control, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 335–344 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By B Cook, M Karakus, J Light, E K. Chanda

The aim of this research is to develop a model through which the appropriate paste recipe parameters, for example solid content and cement content, can be derived given the stope specific requirements such as stope dimensions, required strength and curing time. A large number of samples and uniaxial compressive strength (UCS) tests were conducted at the Angas Zinc Mine (AZM) paste plant and testing facility. Samples were varied in tailings solids content, cement content and sand content. The results of this testing were modelled to determine the UCS relationship to each of the parameters. The output of this modelling was an easy-to-use program that outputs the design mixture given the stope dimensions and required stope backfill-curing time. This allowed flexibility within the mining schedule for curing times of cemented paste backfill stopes. The optimised schedule using the stope specific mix design, when compared to the standard 14-day curing time used previously, resulted in a smoother weekly mine production schedule. For the modelled 12-week period (one quarter) the 14-day curing time would result in an eight-day shutdown of the mill. The cost for each shutdown is in the order of $125 000/d. Due to the unique nature of each mine’s tailings, this exact outcome would not be directly applicable to another paste fill site. However, the process in this report could be followed for another site to achieve a system similar to the output of this project. This would ultimately lead to a more optimised backfill and stoping cycle and a more optimised mine.CITATION:Chanda, E K, Karakus, M, Light, J, Cook, B and Harris, A, 2014. Optimisation of the paste fill cycle at Angas Zinc Mine, Strathalbyn, South Australia, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 259–264 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By P. K. Sollars, M. A. Alspaugh

In May of 1996, Twentymile Coal Co. began Longwall mining in the Foidel Creek Mine's East Mine District. These Longwall panels are some of the longest in the world (5.48 km/18,000 ft) with very difficult and challenging geometry. The conveyance system is 1,800 mm (72 in) wide, has a 4,500 mtph (5,000 stph) capacity and employs multiple tripper drives. The trippers include traditional powering designs as well as unique regeneration and braking designs. Several components were developed specifically for this application including a mobile tail brake. Control algorithms were carefully developed and tested via computer modeling. A world production record was achieved in the third month of operation and over 1 million tons were mined in June 1997.

Jan 1, 1998

CONTENTSThe significance of values givenINTRODUCTIONTHE ORE1. DATA SECURED FROM THE MILLING PLANT AS OPERATING BETWEEN 1938 AND 1942(a) Mill Recoveries(b) Mill Flow Sheet(c) Scheelite-Garnet Association(d) Detailed Examination of Mill Performance during 1942Crushing UnitsCoarse Rolls DischargeFine Rolls DischargeFeed to Huntington MillsFinal Ground ProductProduction of -200-mesh Scheelite SlimeSummary of Information on Crushing and Grinding Circuit(e) Detailed Examination of Mill Performance During 1942Concentrating UnitsLocation and Nature of Scheelite LossesSize Distribution of Scheelite LossesDistributor Spigot ProductsSummary of Information on Concentration Circuit(f) Detailed Examination of Mill Performance During 1942 Gravity Plant Feed2. EXAMINATION OF MINE SAMPLES TAKEN DURING 1942(a) Introduction(b) Nature of the Ore(c) Sampling of Mill Feed(d) Testing Technique(e) Minus 1/8-inch Ores-Naturally Broken Fractions(f) Recoveries from Plus 1/8-inch Fractions-Crushed to Minus 1/8-lnch(g) Probable Recoveries Obtainable from Weathered Ores140 Contour120 Contour90 Contour(h) Examination of an Ore Sample from the 90-foot Adit(i) Run-of-Mine Ore(j) Evidence of Maximum Sizes of Scheelfte Present in Run-of- Mine Ore(k) Summary of Information on Liberation and Recovery from the Ores as Mined in 1942 and from the Mill Samples

Jan 1, 1944

By William Petruk, J. H. Gilles Laflamme, Jean Robitaille, Louis J. Cabri

"A quantitative mineralogical study using ore microscopy, electron microscopy, proton-microprobe, image analysis, and material balances was done on samples from the circuit of a concentrator in Quebec, with particular emphasis on the Cu concentrate. The ore contains many different minerals, of which the principal ones, in decreasing order of abundance, are pyrite, sphalerite, chalcopyrite, bornite, galena, tennantite, and pyrrhotite. Minor to trace amounts of covellite, chalcocite, stannoidite, mawsonite, colusite, wittichenite, miharaite, hessite, petzite, electrum, and an unidentified Cu-Pb-Bi-sulfide also occur.Liberation of pyrite is very high in all size fractions for all samples (ranging from 80-91 %). Liberation of chalcopyrite, bomite, sphalerite, and tennantite is relatively high, on average, in the head .sample and Cu concentrate (55-78% and 66-88%, respectively). Most of the chalcopyrite in the flotation tails is not liberated, whereas about one half of the tennantite and sphalerite are liberated in both the flotation and :final tails. Recoveries calculated for the Cu concentrate (on the basis of assays, mineralogy, and image analysis) are chalcopyrite 72. 5%, bornite 73. 7%, sphalerite 41.0%, tennantite 70.9 %, and galena (together with some trace minerals) 34.2%. It was also calculated that the Cu concentrate contains 10.0 % of the pyrrhotite and i.4 % of the pyrite.Materials balance calculations were done on the five major sulfide minerals to determine the mineralogical distribution of 13 major, minor, and trace elements (as determined by protonmicroprobe analyses). It was determined that tennantite is the principal carrier of Bi (67%), As (90%), and Te (86%) in the Cu concentrate. Therefore, removal oftennantite, which is mostly liberated in the Cu concentrate would significantly reduce these three contaminant elements, but at a loss of9% of the Cu, 2.0% of the Zn, and 0.7% of the Ag. The precious metals (Au, Ag) occur as discrete minerals (electrum, petzite, hessite). Some of the Ag also occurs in solid solution in bornite (23.1% of Ag in Cu concentrate), tennantite, and sphalerite. Four of the major sulfides (chalcopyrite, bomite, sphalerite, tennantite) in the ore also contain measurable concentrations of the following trace elements: Se, Cd, In, Sn, and Sb, occurring mostly as solid solutions. Some of these elements are also concentrated (up to about 18%) in minor and trace minerals such as colusite, stannoidite, and mawsonite."

Jan 1, 1999

By G. C. McLachlan

Two papers dealing with Noranda's milling operations have already been presented. The first1 of these covered the initial metallurgical problems connected with the treatment of the ore, while the second2 described the development of the mill from a 500-ton unit in 1928 to one with a daily capacity of 3000 tons less than six years later. At the time the second paper was written the equipment whirh had been installed to bring the mill capacity up to 3000 tons was being tuned into production and for this reason, though each step had previously been tested by a large-scale pilot operation, it was thought advisable to withhold metallurgical data until any starting-up difficulties had been overcome and anticipated requirements fulfilled. These requirements called for an improvement in gold recovery without any lowering of the previous ratio of concentration, and the pages that follow deal with the methods by which this result was obtained. An idea of the character of Noranda concentrating ore will be gained from an inspection of the following mineralogical analysis of the ore treated by the mill during 1933: MineRal Content Mineral ConteNt Chalcopyrite, per cent..... 6.8 Silica and silicates, per cent............ 18.3 Pyrite, per cent........... 22.1 Gold, native and as telluride, oz. per ton 0.15 Pyrrhotite, per cent....... 51.5 Silver, probably as argentite, oz. per tou 0.35 Magnetite, per cent....... 0.5 Thesr figures, which are typical, show that over 73 per cent of the ore treated consists of pyrite and pyrrhotite. Determirnation of the proportion of unrecovered gold associated with each of these minerals was therefore of primary importance in developing a suitable method for improving gold recovery. As pyrrhotite is magnetic, this information was obtained by making magnetic sorting tests on samples of the mill tailing and screening the material thus segregated. This showed that the magnebtic material finer

Jan 1, 1935

By C Roche, G M. Mudd

The Morobe Province of central Papua New Guinea (PNG) has been mined for gold for nearly a century, although it is only in the past decade that large-scale modern mining commenced. The Hidden Valley gold-silver project began construction in mid-2006 and production by August 2009, and is located in the mountains above Wau in the headwaters of the Watut River. The mine is owned and operated by the Morobe Mining Joint Venture (MMJV), which also owns the Wafi-Golpu project, comprising the Wafi gold deposit and Golpu copper-gold deposit, situated in the Watut River Valley. Together they represent a potential mine of the scale of Ok Tedi or Bougainville or bigger. The MMJV is owned in equal shares by Newcrest Mining from Australia and Harmony Gold from South Africa. Although Hidden Valley was the first large-scale mine in PNG to engineer and build a tailings storage facility, poor environmental management during construction and early operations led to significant erosion of waste rock and sedimentation impacts throughout the ~200 km length of the Watut River. Since then, substantial efforts have been made to improve environmental management, especially waste rock placement and storage and water quality management. While the impacts from the Hidden Valley project appear to be reducing, the saga has heightened concerns by many along the Watut River and across PNG about ongoing impacts from mining.This paper presents the results of an ongoing project in the Morobe Province conducted by the Mineral Policy Institute examining the historical, current and future impacts of mining, including community views, social and environmental impacts and the monitoring and regulation of mining. Overall, there is a clear need to more fully integrate social and environmental issues into life-of-mine planning and go above and beyond regulatory requirements. In this way, some of the lessons learnt – by the community, MMJV and government – can be incorporated before, during and long after mining.CITATION:Mudd, G M and Roche, C, ?2014. Mining in Morobe, Papua New Guinea – impacts, assurance and self-determination, in Proceedings Life-of-Mine 2014 , pp 313–336 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By C. Cortés

Flotation is a highly interactive process where a change or disturbance in any part of the circuit results in different final grades and recoveries. This high variability of the circuit metallurgical performance complicates the diagnosis of the scavenger stage at the end of the line, as copper losses in the final tailings are erratic and not totally a consequence of problems in the scavenger section. The assessment of the scavenger performance requires a more elaborated procedure than that used for others stages of the circuit. This communication describes the application of an approach that combines results of laboratory flotation tests with those of surveys performed at the plant. The approach is based on running surveys in one of the circuit lines to collect samples for determining grade vs. recovery curves for total mass and size-by-size fractions. Full mineralogical and liberation analyses are also considered on some samples of the feed and tailings streams selected to have a range of feed grades. A separate sample of the feed is used to run laboratory flotation tests to determine optimum flotation time. Two additional pieces of information are the effect of operating variables on recovery, determined through a test program based on design of experiment techniques, and operational data on every cell in the line gathered by visual observation. The application of this approach to one of the scavenger lines at Los Colorados concentrator (Minera Escondida, BHP Billiton) is demonstrated.

Jan 1, 2007

By Albert J. Kritscher

The Similkameen Copper Project located near Princeton, British Columbia, is operated by the Similkameen Mining Company Limited, a wholly-owned subsidiary of Newmont Mining Company. The orebodies for this project are located on both sides of the Similkameen River for which the project is named. The project involves the processing of an average of 15,000 STPD of andesitic ore, mineralized largely with Chalcopyrite, at an overall plant availability rating of 90 percent. Project life is nominally 15 years. Design for the plant and its ancillary facilities was undertaken by Bechtel Corporation with procurement and construction management by Canadian Bechtel Limited. Construction began in June 1970 and first ore was introduced into process in March 1972. This paper concerns itself only with the project tailing system. The processing plant is located on a bench of glacial stream gravels which overlooks the river canyon. This site represented the only relatively flat area of sufficient size to accomodate processing facilities near to the ore- bodies. Reference is made to Figs. 1 and 2 showing the relationship of project features in plan and profile.

Jan 1, 1973

By Michel Rodrigue, Sylvain Duchesne, Louis Racine

"LocationThe Louvicourt mine is located 20 km east of Val d'Or in Abitibi, in the northwest part of the province of Quebec. The mine is located on Highway 117 connecting Val d'Or to Montreal and very close to the main railway networkHistoryFive successive exploration attempts by Aur Resources resulted in the discovery of the large Louvicourt orebody in June 1989. Construction of the surface and underground infrastructures commenced in the spring of 1991. Milling operations commenced in July 1994. The entire project cost CDN$290 million, almost CDN$20 million less than initially estimated. The concentrator infrastructure cost CDN$64 million.Louvicourt mine is owned in partnership by Novicourt Inc. (45%), Teck Corporation (25%), and Aur Resources Inc., the administrator (30%).Brief DescriptionLouvicourt mine is the largest producer of base metals (Cu and Zn) in Abitibi-Temiscamingue. Its average daily production is nearly 4300 t of ore, for an annual production of 200 000 t of copper concentrate (28% Cu) and 33 000 t of zinc concentrate (57% Zn). After only a few weeks of operation, it had attained its objectives for the most part, and operations continue to be improved."

Jan 1, 2000

By C. J. Clayton, A. J. Haynes, D. A. Hickson, B. Thiele

The Meadowbank Gold Project, formerly owned by Cumberland Resources Ltd. and currently owned by Agnico Mines Ltd., is currently being developed approximately 70 km north of Baker Lake, Nunavut, in the eastern Canadian Arctic region. The Meadowbank Gold Project currently consists of a series of gold-bearing deposits within relatively close proximity to one another, which are planned to be mined primarily as a truck-and-shovel open pit operation using conventional gold milling and extraction. The mining plan indicates that approximately 20 million tonnes of ore will be mined over a mine life of eight to ten years. The operation will generate approximately 160 million tonnes of mine waste rock and about 20 million tonnes of tailings. The site is located in an area of continuous permafrost with mean annual average air temperature of -11°C and average annual precipitation in the range of 100 to 200 mm. Lakes cover a large portion of the site, including some of the ore deposits. The location of the site and the physical environment pose challenges to development. This paper presents some of the designs considered to effectively manage and utilize the waste products of mining at the project. Note that volumes, tonnages, construction concepts, and data presented in this paper are accurate for the time of writing and are subject to change and modification as the project progresses, additional modelling is carried out, and additional data are obtained.

Jan 1, 2007

By Pierre Lacoste-Bouchet, Ian Orford, Mike Cooper

"The Brunswick Mine ore requires fine grinding to achieve liberation of minerals. With 25mm slugs as the grinding media in the ball mill, there was a limitation of a P80 of 25 microns on the CuPb regrind circuit. Results from pilot plant testing showed that the separation efficiency could be increased through fine grinding with a P80 of 15 microns.Options considered were the capital expenditure for new fine grinding equipment such as a vertimill or the use of the present ball mill equipment with smaller media. The Wheelabrator Allevard Enterprise (WAE) millpebs product was chosen because it was a smaller media with a distribution of steel pellets between 5 to 12 mm.With millpebs the grinding product was reduced in size from a P80 of 25 microns to a P80 of less than 19 microns. The power draw to achieve the finer grind was 13% lower while the media consumption was 57% higher.The millpebs grinding media was also tested on the zinc regrind mill. The purpose of the trial was to understand what the media consumption and power draw would be with the charge lowered to achieve an equivalent grind as obtained with the slugs. The media consumption was approximately the same and the power draw was reduced by 30%.INTRODUCTIONThe Brunswick Mine ore is a complex Pb-Zn-Cu-Ag sulphide ore with the grain size of the minerals varying from submicron blebs to particles up to several mm in size. The ore is ground progressively finer in stages. The first step of grinding occurs in the SAG mill circuit. The product sizing 80% passing 95 microns is fed to the CuPb prefloat stage (Cormier et al 2002). The tails from the prefloat is then further ground to a P80 of 35 microns for the CuPb rougher flotation. The final liberation for the CuPb circuit takes place in the CuPb regrind circuit. The regrinding mill further liberates the cleaner scavenger concentrate, second cleaner tails and third cleaner tails. The reasoning for the progressive grinding is to achieve fine grinding without having galena over ground due to its high specific gravity that preferentially sends it to the cyclone underflow."

Jan 1, 2006