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By R. Dahlke, J. A. Finch

A campaign was undertaken during the greater part of 2001 at Noranda’s Brunswick mine concentrator to investigate the role of gas (air) superficial rate (Jg) distribution to the cells in the final Zn cleaning stage (bank of seven cells), i.e. the Jg profile. Three profiles were tested: “balanced” (each cell with same Jg), “increasing” (from cell 1 to 7 proportionally), and “decreasing,” along with an “as found” profile in a parallel control bank. The experimental design allowed for two confounding factors, changes in ore type and circulating load. The increasing profile consistently gave the best metallurgy (down-the-bank grade / recovery). The reason for improved performance was the response of the first cells in the bank where the increasing profile gave the highest selectivity of sphalerite against nonsulphide gangue (NSG). The other profiles gave lower selectivity and higher zinc recovery in the first cells, which combined to reduce overall bank performance. The higher selectivity with the Increasing profile is attributed to the low gas rate in the first cells reducing water recovery and hence limiting entrainment of NSG. The increasing profile was subsequently configured for all four stages in the zinc cleaner circuit.

Jan 1, 2004

By P. Maltby

"Smelter Feed Preparation Falconbridge Nickel Mines Limited, Sudbury Operations, operate 4 mills and a smelter to produce Bessemer Ni-Cu matte for refining in Norway. Approximately 1400 tons of Ni-Cu concentrate are produced daily which are dried from a filter cake containing 14% moisture to 7.5% moisture pellets. Concentrate drying is done by 4 rotary dryers as shown in the 'Dryer Performance Data' at four separate locations. The dryers all form part of the feed preparation stage prior to sintering.The rate at which concentrate can be produced through the Sinter Plant largely depends on the quality of the dried product which makes up the feed. A feed consisting of mostly hardened pellets will sinter much faster than a feed which consists of broken pellets and dust. Dryer products require:1. Good green strength to survive shipment and handling.2. Correct moisture content.3. Sizing within a range ideally from 1/4"" to 10 mesh in diameter.Dryer performance data is shown in Table 1. Capital costs between the two types of dryers are comparable with the Roto-Louvre producing generally a larger better-formed pellet. Dryer feeds differ in that at both Hardy and Fecunis, Ni-Cu concentrate filter cake is the only feed while at the Falconbridge slurry plant the dryer feed can consist of up to 10% by weight of smelter flue dust resulting in a combined concentrate-dust pellet for feeding to the Sinter Plant. To the Falconbridge Sinter Plant dryer, dust and limestone are added for the smelting process in addition to Ni-Cu concentrate."

Jan 1, 1972

The concept and practices of co-extraction of coal and methane have emerged in recent years to improve coal mining safety and productivity, utilise methane resources, and reduce fugitive emissions in environments that involve multiple coal seams with low gas permeability. The system of co-extraction integrates two previously separate operations of coal extraction and methane drainage. The coal mining operation reduces the coal seam gas pressure and creates strata fractures that enhance gas migration and capture from mine workings and surrounding coal seams. An effective methane drainage operation produces a consistent and quality clean energy product and, at the same time, reduces the methane concentration in underground workings and the methane content and pressure in the adjacent coal seams. This is a significant benefit in preventing gas explosions and outbursts, and promotes a safer and more productive coal mining environment. In addition, the co-extraction with enhanced utilisation of coal mine methane directly leads to a reduction in greenhouse gas emissions. Presented in this paper are key results and outcomes of a major Asia Pacific Partnership project carried out from 2008?2012 by the Commonwealth Scientific and Industrial Research Organisation in collaboration with Huainan Mining Industry (Group) Co. Ltd., to develop and demonstrate advanced technologies at Pansan mine, Huainan, that promote the concept and practices of co-extraction of coal and methane. A comprehensive approach was adopted to assess fundamental strata and coal seam methane conditions during mining, and optimise mining sequences and coal mine methane capture systems. The project has not only demonstrated a practical approach and procedures of advanced co-extraction of coal and methane systems but also delivered significant benefits and outcomes in improved mining safety and efficiency, increased methane capture and utilisation, and reduced fugitive methane emissions at the Pansan mine. The total mine gas drained increased from 25.36 Mm3 in 2007 to 47.0 Mm3 in 2011, and the mine gas utilisation increased from 4.94 Mm3 in 2007 to 21.15 Mm3 in 2011. The improved methane capture resulted in reduced coal and gas outburst risks and methane concentration levels at the mine workings, provided a safe working environment for coal production, boosted the mining efficiency, and eliminated methane related incidents over the last few years at Pansan mine.

Aug 1, 2013

By John A. Dresser

At a meeting of the Montreal Branch of this Institute in the autumn of 1928, the Honorable J. E. Perrault, Minister of Mines for Quebec, announced his intention to enlarge the Bureau of Mines of Quebec by the addition of a Division of Geology. The announcement further stated that the geological work proposed should be begun in the following field season, and that it should be on a substantial scale of not less than six field parties. The announcement was received with enthusiasm and much encouragement was held out for the success of the effort. It has been thought that the present meeting would be a suitable occasion to present a report of progress of this Division for the three years that have since elapsed. History This, of course, is not the first geological work to be undertaken by the Bureau of Mines. Prior to the establishment of the Division of Geology, the Bureau of Mines had not been unmindful of the need of geological research, but it had always lacked the means necessary to any sustained work. For more than 25 years the late J. Obalski was the Superintendent and the sole occupant of the Bureau. Yet, even then, Obalski, besides routine visits to mining districts, made a few exploratory trips along the north shore of the St. Lawrence river and one to Chibougamau, but without the equipment necessary to effective results.

Jan 1, 1932

By Z. X. Cui, X. D. Ma, L. Contreras, X. Jiang

The bath smelting process is one of the key techniques applied in the current copper smelting production, which includes bottom-blown and side-blown furnaces differently from the direction of gas injection. Liquid matte and slag form the bath in the smelting furnace. To better understand the fluid dynamic of the bath and further optimise the operational efficiencies in the smelting furnace, physical simulation of the bath in the horizontal cylindrical reactor was investigated. The mixing behaviour and wave motion of the bath in the bottom-blown and side-blown furnaces have been studied systematically. The mixing time is reduced with increasing gas flowrate and height of the lower layer, while it increases significantly with increasing the height of the upper layer. The wave motion in the furnaces can be suppressed with the presence of the upper layer, but the bath with a lower density generates higher wave amplitude. Overall, in comparison to bottom-blown mixing, side-blown mixing shows better mixing efficiency and lower wave amplitude. In addition, the plume eye and splash phenomena occurred in the bottom-blown furnace have also been studied. INTRODUCTION The bath smelting process is one of the principal techniques used in the current copper making industry. The bath smelting techniques have some advantages of high smelting efficiencies, low capital costs, low energy consumptions and reduced dust generations. From the direction of gas injection, the bath smelting techniques can be classified into five general types: (1) top-submerged blown, e.g. Ausmelt and Isasmelt; (2) top-suspended blown, such as Mitsubishi smelting; (3) top direct-to-copper smelting, such as Outokumpu smelting; (4) submerged side-blown, such as Noranda and Teniente smelting; (5) bottom-blown, such as Dongying Fangyuan bottom-blown copper smelting furnace (Schlesinger et al., 2002; Zhao, Cui, & Wang, 2013; Perez-Tello et al., 2018). In many metallurgical processes involving bath smelting situations, the objectives of agitation are to accelerate the bath homogenization inside the bath which makes a significant contribution to process efficiency and product quality. Extensive mixing studies have been performed for ladle refining and tundish associated to steelmaking processes through water physical models. Mixing time is the parameter used to quantify the efficiency of mixing in the liquid bath and the first empirical equation predicting mixing time was proposed by Nakanishi, Fujii and Szekely (1975). Later on, more systematic studies have been carried out on ladle mixing and researchers realized that the geometry of the vessel could also have a significant impact on mixing (Khajaviand & Barati, 2010). However, the investigations using ladles which are vertical cylindrical vessels, cannot be directly used in the study of mixing behaviour in horizontal cylindrical reactors. Understanding the fluid dynamics inside horizontal cylindrical bath smelter is useful from the industrial perspective. Recently fluid dynamics of horizontal cylindrical smelter related to copper smelting have been studied by the authors’ group (Shui et al., 2015a, 2015b, 2018a; 2018b; Xu et al., 2018, 2019).

Jan 1, 2019

By A. Alfantazi, M. Mahon

The current efficiency in zinc electrowinning is a key indicator of the performance in a cell house. This is typically measured by comparing the weight of the cathode to the electric current applied over the deposition period. This calculation can be cumbersome because the deposition history of each pulled set of cathodes will be unique. It will depend on the applied current and electrolyte conditions from onset of deposition until the cathode is removed and stripped. Not accounting for these can introduce errors to the calculated current efficiency. To track this history and reconcile the current efficiency, a time based computer process simulation was made of an industrial cell house. Two months of hourly recorded measurements from the cell house were input to the simulation, including current to the cell houses, concentrations, flow rate, and temperature of the purified leach solution. The simulation’s predictions of the concentration of zinc in the circulated, spent, and returned electrolyte all showed good agreement with the industrial data over the two months giving a validated zinc balance. The simulation was then tuned to predict the instantaneous current efficiency for all times, as well as the expected masses for the deposits given their deposition history.

Jan 1, 2014

By A. J. Smethurst, S. B. Hearn, S. Boskovic

"The problem of crud in solvent extraction is well known and can be a major handicap and operational issue. Colloidal silica can cause a number of problems including interfering with flocculation, initiating crud formation in solvent extraction (SX), decreasing SX extraction kinetics and increasing SX phase disengagement times. Flocculation can be utilized to remove some of the troublesome particulates but colloidal silica does not flocculate well and presents a different problem requiring innovative chemistry to solve. Coagulants can assist by facilitating the removal of this colloidal silica by promoting the agglomeration of the particles to a size which can be more readily settled or flocculated in solid/liquid separations. This is achieved when the coagulants bind to the surfaces, displacing water and destabilizing the colloids, whereupon they bind to each other which facilitate their amenability to conventional flocculation.INTRODUCTIONSilicate minerals are abundant in the upper layers of the Earth’s crust, and there are few mineral deposits without significant levels of gangue silicates. Silicate minerals include the pure silicas, such as quartz and other silicate containing minerals such as clays, chlorites, feldspars, as well a vast number of less common silicates.During the acid leaching of ores, two types of fine silica particles are released in solution from the dissolution of the host mineral. The first consists of partly depleted silicates, and the second is colloidal silica, which forms in solution. It is this second type which has the potential to cause crud formation as well as poor solution clarities, slow filter rates and excessive flocculant usage (Figure1)."

Jan 1, 2016

By Gui Hua Wang, Xin Chang, Zhong Ying Han, Yuan Fang Cheng

"According to the geometrical characteristics of triangular meshes, Triangular Mesh Split Model (TMSM) based on a finite element method is proposed to simulate cracks propagation of fractured rock. With TMSM, stress intensity factors at the tip of cracks are obtained by far-field contour integral. Meanwhile, vector cross multiplication is adopted to fix the cracking unit, and maximum circumferential stress criterion is used to determine the crack propagation angle. Then crack dynamic propagation simulation is realized through local mesh split or node movement at crack tip. With this method, the crack can split apart a unit or extend along the mesh boundary, so the crack could extend along an arbitrary path and is not restricted by the initial mesh. The applicability of the mesh split algorithm is verified by the crack propagation examples of wing crack and multi-fractured rock mass with holes. The results show that this method not only has higher accuracy of fracture parameters calculation, but also can accurately simulate the crack initiation and propagation of rock under complex stress state such as tension, compression and shear. In addition, the far contour integral is introduced to determine the crack tip stress intensity factor and there is no need to reconstruct a singular element at the crack tip or to intensify the meshes. Compared with the existing mesh reconstruction algorithm, this method significantly improves the computational efficiency.INTRODUCTIONGenerally, rock is neither the full continuum, nor entirely discrete body, but rich in geological defects, such as joints and cracks. A large amount of geotechnical engineering practice shows that most fracture of rock is closely related to the internal joints, cracks and crack expansion (Horii et al., 1985; Wong et al., 1998). Although the classical theory of fracture mechanics established the theoretical foundation of crack initiation, propagation, so far it is still extremely difficult to characterize the propagation of airfoil-typed crack and the interference problem of multiple cracks. Up to now, the micromechanics mechanism research of crack propagation of fractured rocks has still been a hot research topic of rock fracture mechanics."

Jan 1, 2015

By C. Tong, B. Hilscher

"Sensor based ore sorting is being rapidly adopted by the mining industry. The ability to boost mill feed grade by rejecting waste grade rocks early in the process has proven revolutionary for some operations.This paper briefly reviews the current position of ore sorting in the industry, as well as abridged results from specific operations such as Barrick, McEwen, New Gold, and two others. The paper also introduces the concept of “Profit Recovery™”, the measuring of recovery based on rock profitability rather than just metal content.IntroductionSensor based ore sorting is being rapidly adopted by the mining industry. The ability to increase mill feed grade by rejecting waste grade rocks early in the processing stage is proving to be revolutionary for some operations. These potential outcomes have been too invaluable to be overlooked by both producing operations and prospective projects, thus leading to a wide range of studies on ore sorting. In addition, the advent of ore sorting has also called for a new standard of recovery analysis. The answer to this is the creation of a new term and concept - Profit Recovery™.This paper evaluates the most recent results of ore sorting in the mining industry, and presents the results from five different mining operations. Additionally, this paper will discuss and evaluate the Profit Recovery™ of each of these operations. The goal of this paper is to address common concerns regarding ore sorting rejects and their actual impact on project economics.State Of The ArtOre Sorting is a sensor-based sorting and pre-concentration technology implemented before the processing stage that is used to reduce the amount of waste reporting to the mill feed. This technology can also be utilized to recover ore from previously uneconomic waste, as well as reduce the mill energy and reagent consumption (Tong, 2012). In greenfield projects, sensor-based sorting can further provide value by reducing processing capital costs. The decline of available high grade orebodies and decreasing head grades has further increased the attractiveness of pre-concentration, and subsequently ore sorting (Lessard et al., 2014).There are currently two notable types of sensor-based sorting: bulk and rock-by-rock. Bulk sorting consists of live, high tonnage sorting that can be installed on a scoop or on a conveyor belt with a diverter. Rock-by-rock is a selective sorting method that scans and sorts individual rocks on a conveyor belt, and rejects rocks using air jets or mechanical levers. Figure 1 shows a diagram for a typical air jet sorter."

Jan 1, 2018

By K. V. Konigsmann

"The mill at Mattagami Lake Mines treats 3,850 tons per day. The mill feed is a heavy sulphide ore. It is composed ofPyrite & Pyrrhotite - 40%Marmatite - 17%Chalcopyrite - 2%Magnetite - 10%Talc and other Silicates - 31%The sulphides are intimately associated with each other and often finely disseminated throughout the siliceous gangue. An abundance of middling particles exist in the circuit and a delicate balance has to be maintained between concentrate grades and metal recoveries.The metal content of the mill feed fluctuates between 0.4 to 0.7% for copper and 6 to 13% for zinc.The flow sheet of the concentrator is shown in Figure 1.Eight sample streams have been selected for on-stream analysis, namely1. Copper Concentrate 2. Copper Cln Tailings 3. Copper Tailings 4. Zinc Concentrate5. Zinc Rgh Concentrate6. Zinc Cln Tailings7. Zinc Rgh Tailings8. Flotation TailingsThe mill feed is not being analysed. Only a very coarse sample could be obtained which would be difficult to pump, and the fluctuations in the copper content are too small in absolute terms to have much of a bearing on circuit operations."

Jan 1, 1969

By Peter Rubie

For many years, optimising the relining of hard rock grinding mills has been complicated by: ?? the large number of variables; ?? the low frequency of relines (slow feedback loop); ?? the potential economic risk associated with some changes. A technology that has its roots in the automotive manufacturing industry is being utilised to construct a statistically accurate, configurable, and dynamic simulation of the mill relining process. The effects of various mechanisation and process changes can be rapidly modelled and compared. This technology will provide mill maintenance managers and contract relining companies with the opportunity to benchmark current performance against world?s best practice. It will also provide an opportunity for concentrator plant designers and mill manufacturers to plan for maximum plant efficiency, by designing an infrastructure and equipment combination that maximises annual plant availability. This simulation tool and its associated data set, has been developed using recent grinding mill performance data from mills of varying dimensions, liner configurations, mill relining equipment, operator and tool variables. The simulation will provide an output that the various stakeholders can utilise to make optimal decisions within the given financial, physical and technological constraints of the individual concentrator plant.

Jan 1, 2012

By A. Vien, A. J. Neale, B. C. Flintoff, R. P. Edwards

Froth flotation is one of the most widely used and effective processes for mineral beneficiation. First introduced in 1911, flotation technology has seen many developments over the past 80+ years. Innovations in reagent chemistry, equipment design, process analysis and process control have given design and plant engineers a powerful set of "tools" for process optimization. However, it does not follow that all flotation systems are optimized. As Klimpel observes: "The problem with the industrial flotation process is that it is relatively easy to get reasonable results, and relatively difficult to get excellent results" Klimpel (I 988). While reasonable was sufficient in the past, competitive forces will make excellent necessary in the future. Judging by the activities in many Canadian plants this transition is already underway. This paper examines the role of some of the optimization "tools" in effecting this change.

Jan 1, 1994

By M. Mahdi Aboutalebi, Roderick I. L. Guthrie, Mihaiela Isac

"A preliminary computational fluid flow model has been developed to simulate the flow field of liquid steel in the mold region of billet mold casting. This study reports the effect of different Submerged Entry Nozzles (SEN) on the fluid flow in round and square billet, curved mold, casters. In this study, ANSYS Fluent 14.5 software was used to model the 3-D turbulent flow of liquid steel. The standard k-0 model was used to simulate turbulence in the mold region. Similarly, the species transport equations were discretized to simulate the dye injection test into the mold cavity. The behavior of flow in the mold region was validated against previous experimental work. Predicted results showed that different nozzle types plus the slight curvatures in the molds can significantly influence the flow patterns generated within the mold region, and thereby affect the finished quality of the billet products.INTRODUCTIONConventional Continuous Casting (C.C) processes are presently used to solidify more than 95% of the 1.5 billion tons of steel now produced in the world annually (World Steel in Figure, 2014). This overwhelming contribution of C.C processes for steel production is the prime motivation for this study, in order to understand and optimize the casting processes and thereby reduce internal defects and minimize energy consumption. The primary objective of all C.C processes is to maintain liquid metal purity during the casting process and to create a uniform flow within the caster so as to avoid internal and external defects. In this regard, studying fluid flows in the mold region of continuous caster has attracted many research groups, in both academia and industry, in recent years. The mold region is one of the most critical parts of the C.C. process, in which the transport phenomena associated with this region can significantly affect the quality of the final products. Thus, proper control of fluid flows within this region can be expected to greatly decrease both internal and surface defects such as Oscillation Marks (OM), inclusion entrapment, and Mold Powder Entrapment (MPE). In the casting process, the fluid flow in the mold region can be affected by a number of variables, such as; the flow control system (stopper rod or slide gate), nozzle geometry, SEN (Submerged Entry Nozzle) depth, casting speed, mold dimensions, gas injection, and the application of electromagnetic fields. One of the most influential and easily-changed fluid flows are the jets of steel issuing from the SEN. Nozzle geometries are variable, consisting of a number of exit ports, different port angles, and/or port sizes, the bottom shape of the nozzle, and the SEN immersion depth (Thomas, 2001). The SEN characteristics can markedly affect meniscus surface velocities, surface level profiles, and turbulence levels that can vary in terms of the frequency and the magnitude of their fluctuations."

Jan 1, 2016

By H. K. Fredrickson

"Midwest Joint Venture (MJV) owns a high-grade uranium deposit in northern Saskatchewan. The deposit is located too deep below surface to be mined economically by open pit methods, and as a consequence, present plans are that it will be mined by underground methods. High-grade uranium ore of the type at MJV, encased in weak, highly altered ground and with radon-rich water inflows, has not before been mined by underground methods.The test mining phase of the project, completed in 1989, had three objectives: 1. to evaluate radiation protection requirements associated with the handling of large quantities of radon-rich water and mining highgrade uranium ore in an underground environment; 2. to investigate the quantity and quality of water inflows into the mine; and 3. to investigate ground conditions in and around the ore zone as an aid in determining the production mining method to be used.With information gained from the test mining project, a mining method for the production mine has been devised. Level plans have been drawn up, ventilation system designed, pumping arrangements made and methods of ore handling considered. All this is to be done in a manner that will be safe for those doing the work underground. Some of the mining methods planned are felt to be unique in that they are designed to cope with mining problems not known to have been encountered before. New problems underground have required new methods to handle them. Remote drilling, blasting, mucking and backfilling form the basis of the planned mining method."

Jan 1, 1992

By Frank Mucciardi, Louca Gourtsoyannis

"Low and inconsistent recoveries of aluminum in the production of aluminum-killed steels triggered an extensive research program on the aluminum dissolution kinetics at McGill University. This research program led to the understanding of the melting kinetics and the hydrodynamics of making additions to liquid steel. It showed the necessity of gradually releasing liquid aluminum at the bottom of the ladle to increase aluminum recoveries.This knowledge was verified in practice at Sidbec-Dosco, where a prototype aluminum wire injector was developed and installed for the first time in Canada. The industrial application showed the validity of the theoretical and experimental results and provided further data for the mathematical modelling of the process.The development of the theoretical understanding, the industrial results, and the interaction between university and industry are presented.IntroductionAluminum is an extremely effective deoxidizer of molten steel; nevertheless, an efficient method of adding it to steel has eluded steelmakers for many years. It is well known that the recovery of aluminum used in the production of aluminum killed steel is both low and eratic, especially when the aluminum is added in the form of bars or lumps from chutes into filling ladles, a practice still common in many of today's steelmaking operations.As the density of aluminum is only about 40010 that of molten steel, it is conceivable that loose bars of aluminum simply ""thrown"" into a filling ladle will resurface before any appreciable dissolution of the aluminum has occurred, particularly if the depth of penetration of the additions is small. Before one can conclude that this is indeed the case, it is necessary to know the melting kinetics of an aluminum addition. Back in 1972, this information was not available; hence an extensive research program which focused on melting kinetics was initiated at McGill University."

Jan 1, 1982

By F. Rosenblum, P. Spira

A laboratory test has been developed for assessing the patential hazard from self-heating of sulphides. The technique consists in measuring the rate of self-heating, from near ambient temperature, of a 1-kg sample in an insulated chamber controlled at standard atmospheric conditions. The rate of oxygen consumption is monitored simultaneously. The paper describes the apparatus and test procedure, and reports data on various ores and mill products. Tailings known to have caused problems underground when used as backfill have shown heating rates of over 3°C/h. A preliminary hazard rating guide for backfill based on the known behaviour of several sulphide materials is presented.

Jan 1, 1981

By D. A. Sloan

THE Young Davidson mine is of particular interest because of the extremely low costs and also because of the novel method devised for pillar recovery. The outstanding feature of the mining system is its simplicity and regularity. The successful operation of this property was de-finitely assured by the rise in price of gold in 1934. The mine is situated in the southeastern part of Powell township, about forty-five miles southeast of Timmins, Ontario. The local town, two miles distant, in Matachewan, which is twenty-five miles northwest of Elk Lake, the terminus of a branch line of the Temiskaming and Northern Ontario railway. Mining is confined to one of a group of five claims. The ore occurs in a body of syen.ite porphyry which is intruded along a contact of Timiskaming conglomerates with Keewatin altered ruffs and greenstones. This intrusion extends in a north-easterly direction from Davidson creek for 3,300 feet, of which length 700 feet lies within the Matachewan Consolidated property. The maximum width of the mass is 700 feet; it tapers off as it extends east-ward. The Orebody.-The syenite porphyry is of three types, probably all differentiates of the same magma, but only one of them is gold bearing. This is a cherry-red, coarsely crystalline type which forms the central part of the mass. It is more brittle and more fractured than the others, and consequently has allowed easier access to the gold-bearing solutions. The gold has been brought in by a branching system of quartz veins and along minute fractures. It is mainly in the form of very fine 'free' gold, but also occurs in pyrite, which is the chief metallic mineral.

Jan 1, 1942

By G. R. Morin, A. R. Leggett

"In 2010, closure of the Kidd metallurgical plants along with structural changes to the historical ratios of payable metals led to a review of the operating targets at the Kidd concentrator. Previous operating data indicated that there was potential for significant revenue improvements by lowering the copper concentrate grade. An extended plant trial was carried out in 2011 to quantify the revenue impact of lowering the copper concentrate grade from 27% to 25% Cu, as well as identifying the lower practical grade limit. Analysis of plant trial results indicated that the process change increased net concentrate revenue by $2 M to $3 M per year at current metal prices. Increased revenue was primarily associated with higher silver recovery to the copper concentrate while the marginally improved copper recovery was essentially offset by increased zinc losses and concentrate treatment and freight charges.Details of the plant test work results are discussed in this paper along with the financial analysis of the revenue improvements. Key areas covered are major revenue components of the concentrate contracts, validation of revenue improvements through plant testing and sensitivity of the improvements to market and plant variables.INTRODUCTIONThe Xstrata Copper Kidd concentrator commenced operations in 1966 treating ore from the nearby Kidd volcanogenic massive sulphide deposit in Timmins, Ontario (Thompson, Amsden, & Chapman, 1978). Over the years, up to four separate grinding-flotation divisions were in operation with throughput rates peaking at 4.5 million tonnes per year of ore in the mid 1980’s (Kelebek & Francis, 1998). As the mine production decreased and grinding throughput increased operations were gradually cut back and since 1998 two grinding-flotation divisions have been in operation treating Kidd mine ore. One of the idled divisions has been intermittently used for custom milling operations, most recently the Ni/Cu Montcalm deposit, which was milled from 2004 to 2009. Current budget throughput rates, limited by mine production, are 2.4 million tonnes per year of ore grading 1.7% Cu, 4.3% Zn and 55 g/mt Ag. Concentrate production rates are 150,000 tonnes per year of copper concentrate and 155,000 tonnes per year of zinc concentrate. The Kidd mine is the deepest base metal mine in the world with mining operations down to 9600 ft.In 2010 the copper and zinc plants at the Kidd metallurgical site, which previously treated the concentrate production from the Kidd concentrator, ceased operation due to a variety of economic and market factors. Following the metallurgical plant closures the Kidd concentrator products were shipped off site for treatment. The majority of the copper concentrate is now processed at the Xstrata Copper Horne smelter in Rouyn-Noranda, Quebec with the majority of the zinc concentrate processed at Canadian Electrolytic Zinc Ltd. in Valleyfield, Quebec."

Jan 1, 2013

By H. A. Pearse

Introduction This paper may be regarded as another instalment of the story of the Britannia milling operation. From time to time in the past, articles (1) have been prepared by various members of the staff, giving details of crushing, grinding, and flotation treatment for the normal Britannia ore. It is not, therefore, the purpose of the present paper further to discuss what may be termed the standard milling operation, but rather to describe a flotation practice which has been evolved directly as a result of the general depression. The normal Britannia ore mixture contains chalcopyrite and pyrite as the main sulphide minerals, together with minor gold and silver values. The standard milling practice, evolved during several years of experimentation, provided for bulk flotation of the two sulphides after primary grinding to only approximately 35-mesh, followed by regrinding of the bulk concentrate, with subsequent separation of the chalcopyrite and pyrite to make a marketable concentrate of each mineral. However, in the face of a constantly increasing surplus of copper on the world markets and a steadily falling price during 1930 and 1931, it was obviously undesirable to continue producing large quantities of this metal at a loss. On the other hand, Company officials were very reluctant to add to unemployment by closing down entirely, provided operations could be continued without excessive losses. Fortunately, the existence was known of a fair sized body of ore, quite low in copper and relatively high in zinc content, with also a very considerably higher gold value than that prevailing in other sections of the mine.

Jan 1, 1934

By X. L. Song, G. W. Zhang, F. Y. He, X. Xiao, L. L. Huang

This paper presents the design of a model and process simulation of a vertical mill based on a population balance model. The simulator, Simulink, which is a systematic modeling tool within MATLAB, was used for the simulation. The model will simulate the real time variation trend of each size fraction in the feed material to the vertical mill, as well as, the size distribution of the batch grinding products. Breakage distribution and selection functions are calculated based on batch grinding laboratory experiments using Back-calculation method. The simulated size distributions correspond with those in the actual mill operation having a relative error of less than 0.50%. This work lays the foundation for future research on modelling and simulation of a closed circuit grinding operation consisting of a vertical mill with hydrocyclone classifier.

Jan 1, 2016