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"Drilling twinned holes is a traditional technique used for verification of intersections of high-grade mineralization, testing of historic data, or confirmation of drillhole data during geological due diligence studies. Twinned holes can also be used for special tasks such as correcting earlier data that are recognized to be biased. Successful implementation of the twinned-holes technique requires thorough planning. Experience suggests that good practice is to drill twinned holes no more than 5 m apart. Many unsuccessful twinned-holes programs could possibly have failed because the twinned holes have been drilled too far apart.Twinned holes are best compared by mineralization intersections and, if the data and geological characteristics of the deposit permit, by samples, equal-length composites, or geological units. Variables to be verified by twinned holes should include the thicknesses of the geological units of interest (e.g., mineralized thickness) as well as the presence of significant geological features (e.g., ore/mineralization contacts, alteration, etc.).A formal, rigorous analysis of twinned-hole data is essential. Repeatability of sampling, analytical results, and bias must be analyzed and statistically quantified. The number of twinned holes required for conclusive statistical and geostatistical analysis can be as high as 20-30, in particular where the studied variables are characterized by high short-range (local) variability.The efficiency of this approach is demonstrated by several examples of successfully applied twinned-hole projects, including applications to orogenic gold, mineral sands, bauxite, and iron ore deposits."

Jan 1, 2009

By P. Brad Marchant

"To properly evaluate bioleaching as a commercial alternative for sulphide concentrate oxidation, to enhance cyanidation response, it was necessary to define and optimize critical operating variables outside of the laboratory under less than ideal conditions. A complete bioleach testing laboratory was installed at Equity Silver Mines to aid in pilot plant flowsheet development.A continuous pilot scale bioleach circuit was constructed to provide data for meaningful valuation of biotechnology on a plant scale. The important operating and design parameters for pilot plant testing are described. Plant scale feasibility and design considerations, to treat a refractory pyrite/arsenopyrite concentrate produced from flotation tailing, are discussed. 1.0 BACKGROUNDEquity Silver Mines Limited is located 37 km southeast of Houston in central British Columbia. A 70% owned subsidiary of Placer Development Limited, Equity is Canada's largest open pit silver producer. Ore reserves were estimated at 28 million tonnes grading 106 g Ag/t, 0.96 g Au/t, and 0.38% Cu from two pits, the Southern Tail and the Main Zone. Detailed geology and mineralogy has been described previously(1,2).The concentrator flowsheet consists of conventional crushing, grinding, flotation, and dewatering circuits processing 5 300 tonnes per day of new mill feed. The resultant Cu/Ag/Au concentrate is trucked to Prince Rupert for shipment to market. Mill tailing reports to a closed circuit pond from which mill process water is recycled. Details of the concentrator process have been reported( 2,3).Production in the Southern Tail pit, from July 1980 to March 1984, resulted in mill tailing averaging 0.9 g Au/t and slightly more than 20 g Ag/t. Historic data indicate that unrecovered gold is associated with arsenopyrite and silver losses with unliberated tetrahedrite or minor silve,, bearing sulphosalts. Main Zone mill tailing, which commenced production in March 1984, contains approximately 0.5 g Au/t which is probably associated by majority with silver bearing minerals and pyrite while silver losses, 30-40 g Ag/t, are associated with unliberated silver and antimony bearing minerals."

Jan 1, 1985

By Samantha Beeson, Collin Poole

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

Jan 1, 1997

By B. Haugrud, E. McConaghy, M. Collins, G. Smith

"Laterite ore processed by the Ambatovy Joint Venture in Madagascar is primarily limonitic in nature, containing about 40 to 45 wt.% Fe and 1.1 wt.% Ni. The process selected for treatment of this ore considered the behavior and deportment of iron in each step of the commercial flowsheet. Such considerations included nickel and cobalt recoveries, reagent consumptions, liquid-solid separation characteristics, and both intermediate and final product purities. In this paper, the results for the first three years of commercial operation at Ambatovy are compared with the design parameters, with emphasis on the behavior of iron.INTRODUCTION The main component of nickeliferous laterite ore is iron, and the Ambatovy deposit in Madagascar is no exception. The Ambatovy ore is primarily limonitic in nature, containing about 40 to 45 wt.% Fe and 1.1 wt.% Ni. Since the nickel content of the ore is roughly one fortieth that of iron, successful recovery of nickel depends on close control of the behavior of iron in the selected process. Exploration of the Ambatovy area, which is located to the east of Antananarivo, the capital of Madagascar, dates back to 1960, when the Malagasy Service Géologique carried out a limited pitting and auger-drilling program. Phelps Dodge Corporation (now Freeport-McMoRan Copper & Gold, Inc.) was granted the mining rights to the property in 1995 and carried out significant resource definition, environmental assessment and engineering studies between 1995 and 2001 (Chou, Rooke, Williams, & Hanks, 2000). In 2003, Dynatec Corporation, which was formed in 1997 by the merger of Dynatec International Ltd. and Sherritt International Consultants Inc., signed a joint venture agreement with subsidiaries of Phelps Dodge to evaluate the Ambatovy Nickel Project. Dynatec then became extensively involved in the development of the project, including process selection and pilot plant testing in Fort Saskatchewan (Collins et al., 2004, 2005, 2009; Collins & Vardill, 2005). Dynatec acquired 100% of the project in 2005 and a joint venture between Dynatec, Sumitomo Corporation, Korea Resources Corporation (KORES), and SNC-Lavalin Inc. (SLI) was formed shortly thereafter to provide funding for further development. SLI was responsible for engineering, procurement and construction management of the commercial facilities. Sherritt International Corporation reacquired Dynatec Corporation, including 40% ownership in the Ambatovy Nickel Project, in 2007. The current Ambatovy Joint Venture consists of an integrated nickel and cobalt mining, processing, refining and marketing enterprise between subsidiaries of Sherritt (40% owner and mine and plant operator), Sumitomo Corporation (32.5% owner) and KORES (27.5% owner)."

Jan 1, 2016

By D. Ouellet, G. Gagnon, B. Simser, V. Falmagne

The mining sequence at Noranda’s Bell- Allard mine calls for two mining fronts retreating toward a central pillar known as the Retreat Zone between pyramids A and C. It is clear from numerical modelling and rock mass characterization that the induced stress level in the pillar will exceed the rock mass capacity. A mining plan was developed including the redimensioning of stopes in the Retreat Zone, additional ground support and instrumentation in order to ensure the safety of the workforce and to maximize ore recovery while minimizing production delays. The mining strategy consists in allowing the rock mass to fracture and fail progressively while maintaining control of the existing openings with the appropriate support. A systematic analysis of rockbursting and caving hazards was conducted for all accesses and stopes in order to plan ground support and instrumentation needs.

Jan 1, 2004

By M. W. Lewis

A mine's inherent capacity is dictated by factors such as equipment, design, people, processes, and environment. Actual production may approach, but never surpass, the inherent capacity limit. The inherent capacity ceiling can only be increased through re-engineering. Autonomous haulage systems have the potential to significantly increase a mine ’s inherent and realized capacity. Realized capacity may be maximized through the optimization of soft factors such as the processes, behaviours, and organizational systems, controlling asset utilization, availability, and performance. Truck shovel dispatching systems revolutionized the utilization of mobile assets. High-precision global positioning systems (HP-GPS) are increasingly being adopted to optimize equipment performance. Real-time maintenance management systems have the potential to revolutionize equipment availability and significantly reduce cash costs. This paper explores some of the solutions available to mines to increase their inherent capacity and maximize their realized capacity.

Jan 1, 2004

By M. W. Lewis, B. Sambirsky, J. Werner

A mine’s inherent capacity is dictated by factors such as equipment, design, people, processes, and environment. Actual production may approach, but never surpass, the inherent capacity limit. The inherent capacity ceiling can only be increased through re-engineering. Autonomous haulage systems have the potential to significantly increase a mine’s inherent and realized capacity. Realized capacity may be maximized through the optimization of soft factors such as the processes, behaviours, and organizational systems, controlling asset utilization, availability, and performance. Truck shovel dispatching systems revolutionized the utilization of mobile assets. High-precision global positioning systems (HP-GPS) are increasingly being adopted to optimize equipment performance. Real-time maintenance management systems have the potential to revolutionize equipment availability and significantly reduce cash costs. This paper explores some of the solutions available to mines to increase their inherent capacity and maximize their realized capacity.

Jan 1, 2004

By Philip J. Guerney, S. O’ Leary, André R. Laplante

"Concentration of dense mineral phases (s.g. greater than 7) for characterisation of individual particles and mineral grains by electron microscopy is a well established technique for the detection and characterisation of valuable mineral phases. This presentation describes the type of data that can be generated when both the concentration and characterisation steps are carried to higher levels than typical. The concentration step is the gravity recoverable gold (GRG) protocol that has been used for over 100 ore samples. This protocol yields three concentrates that represent three progressive degrees of grinding and liberation. Three ore samples were subjected to the GRG protocol and three size classes of each concentrate were characterised using the mineral liberation analyser (MLA), an automated, EDX-based tool. This presentation briefly presents GRG and MLA technology, and then focuses on the results obtained: photos, images and databank, the latter used to evaluate the extent of particle liberation and particle shape.INTRODUCTIONThe practice of gravity concentration prior to characterization of mineral samples using electron beam microscopy is solidly established. Typically, a dense liquid such as tetrabromoethane (density: 2.97 g/cm3) is used to reject the light phase, typically carbonates and silicates. The heavy phase, mostly sulphides, oxides and precious metals, is then mounted in a polymer matrix and examined using beam microscopy.Both the pre-concentration and the sample preparation/examination steps have seen significant advances over the past ten years, but are not without problems. A major issue of preconcentration is that most dense liquids used are toxic, tetrabromoethane being a prime example. More recently, a dense liquid with little or no toxicity (sodium polytungstate, distributed by Sometu Inc.) has become commercially available, but its use is limited to relatively low s.g. splits around 3 g/cm3. Dense liquids are viscous, and their use is limited to particles coarser than 37 µm, although centrifuge separation can probably lower this limit to around 10 µm. Magnetogravimetric separators can achieve much higher specific density cuts, but upgrading below 100 µm or in the presence of magnetic minerals remains a problem. The use of centrifuge separators such as the 3-in Knelson Concentrator (KC MD3) or the Falcon 4-in SB (SB4) is particularly attractive since these units can easily process large masses (by laboratory standards) and achieve a cut density well above those of dense liquids."

Jan 1, 2003

By S. Banisi, H. R. Iran-Manesh, M. R. Shayestefar, H. Shekarchyan

"Flotation circuit of the Sarcheshmeh copper mine processes 41,000 t/d of ore at 0.9% Cu. Chalcopyrite and chalcocite are the main copper bearing minerals which are concentrated in three stages of roughing, cleaning and recleaning. Due to major changes in the reagent addition regime, make-up balls combination, and throughput to the primary mills an increase of 2.6% (from 87.2% to 89.8%) in the sulfide copper recovery was obtained during a period of six years. In order to investigate a rather large variation of monthly recovery (i.e., from 89.0 to 91.2%) of the plant after implementation of the changes, a study was undertaken to trace mineralogically possible causes of such a behavior. During a five-month period, monthly composite samples of feed and tailings of the flotation circuit were taken. At the first stage, the polished sample preparation method and the mineral counting procedure were optimized through a detailed study. Liberation degrees and types of locking of valuable minerals in both streams were determined in three size intervals of +75, +38 and –38 microns. It was found that for both recovery levels (i.e., 89% and 91.2%), there was no significant difference in the liberation degrees of valuable minerals in the flotation feed; the average liberation degree was 57.1% with a standard deviation of 3.1%. On the other hand, in the case of high recovery level (i.e., 91.2%) there was a reduction of 15% in the amount of free valuable minerals in the tailings stream. The highest reduction was found for the -38 micron size range in particular for chalcocite. It was concluded that compared to the grinding circuit the change in the flotation circuit had the significant effect on the plant recovery increase.INTRODUCTIONThe Sarcheshmeh copper ore body contains 1 billion tonnes averaging 0.90 % copper and 0.03 % molybdenum. It is located southeast of Iran and currently processes 41,000 t/d of ore. The concentrator has been in operation since 1982. Last year, the concentrator produced 130,000 tonnes of copper and 4,000 tonnes of molybednite concentrate. An expansion to the plant which will increase the throughput by 50% is under construction and will be effective in 2003.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 constitute the main component of the plant final tailings. A concentrate with an average grade of 30.5% Cu is obtained after cleaning and re-cleaning stages. At the present time, the average sulfide copper recovery of the plant is 89.8% which varies depending on the operating conditions and ore type."

Jan 1, 2003

By M. Maccagni, J. Nielsen

"Engitec developed the EZINEX® Technology some years ago to recover zinc from oxidized secondary sources. This process works at neutral pH and can take into solution any kind of metal, but iron is rejected and remains in the leach residue. Engitec has also developed the FLUOBORATE System, which is able to recover lead and copper from relevant concentrates. Some iron is leached into solution, but it is controlled through a solution purification bleed treatment. In this paper, we present the two above mentioned technologies. They are very different with regards to the chemical environments and the methods used to control the build-up of iron in solution, but both systems are suitable to treat iron bearing feeds.INTRODUCTIONIn the design of hydrometallurgical processes, the most important issues are the water balance, the filtration and the control of impurities. This is particularly important when the process includes an electrowinning unit. Iron has been one of the most problematic elements to remove, especially in acidic environments.Engitec Technologies has developed several hydrometallurgical processes divided in two different electrolyte lines: the Fluoborate Technology for the recovery of lead and copper from primary and secondary sources and the Ammonium Chloride System based on a chloride electrolyte for the recovery of different non-ferrous metals, among which zinc and lead are the most important. In both cases, iron is present in the feed and needs to be eliminated from the solution. It is not a problem for the quality of the metal produced, but when present in solution, it consumes the basic electrolyte reducing its free concentration which is very important both for both the leaching and electrowinning unit operations. Both process systems are very peculiar, the fluoborate case being much more intriguing."

Jan 1, 2016

By Craig M. Sams

"Variable speed mill popularity has grown during the last 10 years with almost every SAG mill fitted with a variable speed drive. In addition, many of today’s larger ball mills are running a variable speed drive for various reasons. In the past, motor suppliers, especially gearless drive suppliers, have considered mills as being an overpriced rotor. This has lead to mills demanding more torque than what a motor can generate. This paper identifies the reasoning of what has happened and how to mitigate the mismatch in the future. Technical and economical issues are discussed, allowing mill operators to make an informed decision of which path to take in deciding what motor to put on a mill.INTRODUCTIONIt has been generally accepted in the past that motor suppliers and specialists in engineering companies understand the workings of a SAG mill or a ball mill and although there is a fair knowledge within the industry, in regards to mill torque requirements, the advent of variable speed drives in ball mills has shown the writer that a complete understanding of the mill torque requirements are not fully understood.This paper will discuss two aspects of mill torque requirements, (operating torque at various charge levels and start up torque), and how it relates to specifying the proper motor torque requirements.MILL OPERATING TORQUEMill operating torque at a given charge level and density in a mill is essentially constant throughout the entire speed range. For the purpose of this paper, we will take a 38’x 22.5’, 20’ EGL, SAG mill and provide the power draw at variable speeds but fixed loading conditions.The power draw and mill torque of this mill at 8% ball charge and 28% total loading (Highlighted) is shown below as Figure 1, at variable speeds."

Jan 1, 2007

By Jean M. Richardson, Steven R. Williams

"Geometallurgical mapping is a new team-based approach that documents variability within an orebody and quantifies the impact of geology (host rocks, alteration and structure) and mineralogy on grinding, metallurgical response and metal recovery processes. The quantitative, spatially constrained database that results can be readily integrated into 3-D block models and mine planning activities. Thus it is an important tool to reduce the technical risk associated with new mine developments or expansions.Normally undertaken during pre-feasibility or feasibility planning stages, the first step of geometallurgical mapping is to review the geology, mineralogy and other critical parameters and construct a geometallurgical matrix. This matrix provides an objective base to then guide sampling and/or compositing activity for physical property and metallurgical testing. Depending on the project, an array of testing techniques can be used to characterize the ore and feed data into the geometallurgical model.While the range and number of tests needed is project-dependant, the geometallurgical mapping approach allows the development of empirical models or interdependent relationships, grounded by spatially-constrained real data. Extensive data sets collected using less expensive testing methods can be meaningfully correlated with results from complex specialty tests to yield realistic metallurgical recovery data.With geometallurgical mapping completed, the resulting database can be integrated into an overall 3-D block mine model. This means that metallurgical response can be used to realistically forecast the recoveries of specific blocks, perhaps sampled only by drilling. Mine planning can then incorporate the forecasted metallurgical response and generate future project cash-flows; opening the door to economic optimization of the mine exploitation."

Jan 1, 2004

By Mostafa Benzaazoua, Bruno Bussière, Philippe Poirier, Denis Bois

"Environmental desulphurisation is an attractive alternative for the management of acid generating tailings. Non selective froth flotation is the most commonly used method to produce desulphurized tailings, as shown from many previous works (McLaughlin and Stuparyk, 1994; Bussière et al., 1995; 1998; Benzaazoua et al., 2000a,b; 2002). This process can reduce the volume of problematic tailings to manage, by making a sulphide fraction concentrate. The desulphurized tailings can then be used as a moisture retention layer in a cover with capillary barrier effects to prevent acid mine drainage.In this paper, the authors present the results from different laboratory and pilot scale flotation tests which confirm that low sulphide, non-acid generating tailings can be produced by froth flotation. It is also demonstrated that the sulphide concentrates obtained from desulphurisation, when mixed with binders gave good long term results in terms of mechanical behaviour, thus demonstrating that the sulphide fraction can be used for paste backfill material. From previous work (Bussière et al. 1998, 1999) on the evaluation of the operating and capital costs of the desulphurisation process for different ore types, the use of desulphurized tailings to create a dry cover was then studied for several scenarios. The desulphurisation process was compared with other technologies (underwater disposal, engineered soil cover) currently used to prevent acidic drainage (AD) for a new mining operation and the advantages and disadvantages of desulphurisation are presented as well as the costs of this method of AD prevention.INTRODUCTIONThe principal environmental problem of the mining industry today is the generation of acidic drainage (AD) resulting from the natural oxidation of the sulphides in mine tailings. This phenomenon involves the generation of acidity which increase the amounts of toxic metals in the waste water released into the environment by a mine. The environmental impact of the AD has been described in detail in previous work (e.g. Down et Stocks, 1977; Ritcey, 1989; Gray, 1997; Ripley et al., 1996; Aubertin et al., 2002)."

Jan 1, 2004

By T. E. Morimoto

Summary The Research Council of Alberta has made a series of tests with the cyclone washer to investigate the influence of the following operating variables: underflow apex diameter, .specific gravity of medium, pulp ratio, density distribution of the feed, and size analysis of the feed. The results obtained for all the variables studied, except size analysis of the feed, are included in this paper. The tests were made with smaller underflow apex diameters and lower specific gravity .suspensions than have been recommended by Cerchar for the sharpest separations. However, what is considered to be sufficiently good cleaning was accomplished with a suspension of 1.30 specific gravity and an underflow to overflow diameter ratio of 0.56. It was found that the sharpness of separation was not affected appreciably by variations in specific gravity of suspension or underflow apex diameter provided that the density of separation was within the normal range of washery operation. Cyclone performance was not too adversely affected by decreasing the pulp ratio. The minimum figure which gave reasonable separation was 4 to 1. The tests with various density distributions of feed revealed .that the degree of washing difficulty had no detrimental effect on the .sharpness of separation given by the cyclone. Introduction A major problem of the bituminous coal industry today is that of reducing the ash content of the finer sizes (minus 1/4-inch) of coal. This problem is particularly acute in Alberta, where, because of tectonic movement subsequent to deposition, the coals have been rendered friable, and bands of impurities have been intimately mixed with the clean material. As a consequence, most Alberta coals, as mined, contain a large proportion of fines and have a high inherent mineral-matter content. The advent of mechanized mining and stripping operations ?has further aggravated this problem.

Jan 1, 1952

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 J. Coffin, R. Multani, C. Deredin

Glencore’s Strathcona Mill has recently evaluated a pyrrhotite depressant combination of diethylenetriamine (DETA) and sodium metabisulfite (SMBS) to improve its overall plant pyrrhotite- pentlandite (Po-Pn) selectivity, with the aim of producing a higher quality nickel concentrate. In a previous paper, key findings from a comprehensive lab-scale program (completed at the XPS Centre) were summarized where, of the depressants tested, DETA/SMBS combination was found to be the most effective in improving the Po-Pn selectivity. The present paper describes the metallurgical program recently completed at Strathcona Mill; DETA/SMBS evaluation on select plant “grab” samples as well as the major results of three DETA/SMBS plant trials conducted on several plant ore feeds (long-term testing). Interestingly, it was found that there was variable success in using DETA/SMBS depressants on the different plant ore feeds; DETA/SMBS (dosage: 25/50 g/t milled feed) on lower Cu feeds demonstrated a ~2% Ni recovery increase at the same Ni grade. On the other hand, higher Cu feeds demonstrated virtually no benefit from DETA/SMBS at the same plant dosage, suggesting that higher DETA/SMBS dosage (at specific locations) than that tested is required to overcome the increased Cu-DETA complexation for these particular feeds. Glencore Nickel – Sudbury Integrated Nickel Operations owns and operates Strathcona Mill which has been processing Ni-Cu ores from the Sudbury Basin since 1968. The valuable sulfides are pentlandite (Pn) – (Ni,Fe)9S8 and chalcopyrite (Cp) – CuFeS2, majority of the sulfide gangue is pyrrhotite (Po) – Fe(1-x)S [0 < x ≤ 0.125] which is present in two forms: hexagonal Po (Fe9S10) and monoclinic Po (Fe7S8) (Multani & Waters, 2017; Lawson et al., 2014). Currently Strathcona Mill processes ore mined from two primary sources from the Sudbury area and produces two concentrate streams: (1) nickel-copper concentrate that is sent to the Sudbury Smelter; and (2) copper concentrate that is shipped to the Fonderie Horne (Glencore Copper, Rouyn-Noranda).

Jan 1, 2019

By Brad Simser

The Brunswick No. 12 Mine is a 10 000 tonne per day producer of lead-zinc-copper-silver ore that has been in operation since 1964. Mining of this multiple lens deposit is primarily by blasthole stoping techniques between 400m and 1200m below surface. The combination of a high overall extraction ratio and high horizontal stresses (twice the vertical) has elevated the induced stress regime to levels comparable to much deeper operations. Improved mine design, and a better understanding of how the rock mass responds to mining in high stress environments is essential to improve overall recoveries and maintain safety. Seismic monitoring has proven to be an essential tool at Brunswick, for both short-term safety of workers, and long-term strategic planning. Practical case studies illustrate how seismic data, particularly source parameter information, can be used to measure the rockmass response to mining and improve mining layouts in difficult regions. As the global trend continues towards mining in deeper, higher stress environments (e.g. South African Witswatersrand Basin, Sudbury Basin) and as local extraction and thus ground stresses continue to increase in existing deposits, seismic monitoring will play a critical role in mining safely and economically.

May 1, 2003

By Egbert Burchardt

The challenge of "ore bodies getting steadily harder and lower in grade" was brought to light during the first oil crisis in the 70&apos;s. Shortly afterwards, HPGRs appeared on the scene with the promise of saving energy. This combined with developments in wear protection for the rolls ushered in a period of trade-off studies that considered the HPGR as an option to improve project economics or to even make projects feasible at all. Since then, applications of HPGRs at Cerro Verde (Freeport), PTFI Grasberg (Freeport), Mogalakwena (Anglo Platinum), Tarkwa (Goldfields), Boddington (Newmont), Penasquito (Goldcorp), and others have contributed to the creation of a formidable basket of operating experience. Collected through years of successful operation, this information has been used to validate design criteria, scale-up factors and establishing a clearer picture of what "to do" and what "not to do". This paper compares results from various mining operations using HPGRs in their processing plants. Some of the lessons learnt in different operating environments are reviewed in order to support proper implementation and operation of HPGR&apos;s. Some of the development programs that have been initiated to improve wear life of the tyres and to reduce overall operating costs are reviewed. It is timely to summarize today?s experience and analyse its implications for existing and new operations.

Jan 1, 2012

By T. J. Yeomans

"The Kidd Creek metallurgical site is situated twenty-six kilometres east of Timmins, Ontario. The site consists of a 4.5 niillion tonne per year concentrator, a 127,000 tonne per year electrolytic zinc plant and a 79,000 tonne per year copper smelter/refinery complex.The concentrator commenced operations in 1966 and was expanded to its present capacity in 1978. The zinc plant commenced operations in 1972, and the copper smelter/refinery in 1983.The concentrator tailings, and both liquid and solid waste products from the other plants are disposed of within a 1215 hectare tailings disposal area. Here, a method known as centraljzecl thickened discharge is used to form a stable cone-shaped hill. A low permanent perimeter dyke structure encloses the cone and directs cone run-off and thickener overflow to a treatment pond. After treatment with lime slurry, this neutralized effluent is recycled to the metsite and also discharged into the Porcupine River system.The tailings system has been in operation in its present form since 1972, upon completion of dykes and a 110 metre thickener unit. Underflow densities were continuously increased until thickener overloading became a serious problem. This necessitated the construction of a second 110 metre thickener, which was completed in 1983. Target densities have been achieved without difficulty since then.This paper discusses the operation of the tailings system, reviews past performance, and outlines future operations."

Jan 1, 1986

By Fidelis T. Suorineni

The Ashanti mine, located in Obusai, Ghana, West Africa, has been in production for over 100 years, and is currently the second largest single gold producer in Africa. The mine covers a concession area of 434 square kilometers with four main ore types: the quartz vein type, disseminated sulphide type, supergene sulphides and transition ore, and granitoids stock works (visible gold and quartz veinlets). Various mining strategies and methods have been employed for the mining of the narrow-vein ore at Ashanti. In recent years, the mine owner: Ashanti Goldfields Company Ltd. has embarked on a mine-wide expansion program to boost production to a target of one million ounces per annum. With the rapid depletion of its high-grade ore, a suitable mining method was required to produce a high tonnage at low cost to meet production objectives. This paper discusses how the Ashanti mine applied Canada experiences in mining technology to boost production, and the impact that the differing geological settings between the two countries had on the stope performance at the Ashanti operations. Experience developed at the Ashanti mine led to innovative modifications for stope layout and design methods in relatively poor ground conditions.

Oct 1, 2001