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By C. Bazin

"Discrete transfer functions which use the grinding circuit feed tonnage to predict mass flowrates at reagent addition points in the flotation circuit were calibrated using a dynamic grinding/flotation simulator. Predicted values are multiplied by g/t ratios to calibrate reagent flow. The transfer functions are currently integrated into the distributed control system at Les Mines Selbaie.INTRODUCTIONThe development of adequate tools for the control of flotation processes is a continuous challenge. Since the early 1970's there have been numerous publications dealing with practical applications of flotation process control. A review of that literature shows that basic feedback and feedforward controls are the two preferred strategies for flotation processes. Some advanced control schemes started to emerge (e.g. Zaragoza and Herbst, 1988). Feedforward and feedback aim at reducing the effect of the following flotation process disturbances (e.g. McKee, 1991):0 Changes in throughput;0 Changes in feed grade;0 Changes in particle floatability (size, liberation, surface oxidation ... );0 Deliberate changes of set-point;Feedforward strategy acts on manipulated variables, mainly reagent addition, to attenuate the effect of disturbances such as changes of feed rate and grade. The feedforward strategy is driven by measurements of disturbances. Feedback control acts to compensate unmeasured disturbances, such as ore floatability, and to bring process variables to targeted set-points of concentrate or tails grades. The feedback control action is driven by set-point error, caused by either a deliberate change of set-point or unaccounted disturbances.This paper proposes a way to improve feedforward control schemes by addition of a simple model of solids transportation. The feedforward strategy considered in this paper is of ratio type, i.e. the manipulated variable is changed according to a predefined ratio between this variable and the measured disturbances. A typical example is the ratioed reagent addition to mass flow of ore."

Jan 1, 1993

By A. Khan, C. Ciriello, W. Oliveira

"The purpose of this paper is to demonstrate the value of Operational Readiness (OR) in the successful delivery of capital projects. OR can be captured using several technical and practical definitions:• To ensure that the capital project being delivered is capable of an effective start up, ramp-up as planned and a sustainable productive operation.• To function as intended when started, with the appropriate people able to run it.• To increase the value in delivering the product/service through established systems, processes and procedures, while simultaneously minimize risk to both the project and the business.Historically, mining and metallurgical capital projects concentrate efforts on cost and scheduling during the implementation and design phases. The completion and subsequent handover of the operational unit is the main objective.When reviewing the mining and metallurgical industry’s capital projects, it seems many poor results and errors in terms of project design and ramp-up are self-inflicted, with the project neglecting to account for key aspects of the future operational plant. Interestingly, if the problems for the future plant occur during project implementation, they can be prevented during the project phase. It is required to connect the operational knowledge with the project in order to ensure an effective start up and operation.The benefits of OR will be shown to be applicable not only to mining capital projects but also to sustaining and business transformation projects. The paper also includes industrial examples showing both proper and improper use of OR techniques.The paper will cover 3 main areas of Operational Readiness: the OR model, OR related risk elements (i.e., risks associated specifically with the improper execution of OR) and conclude with OR specific case studies."

Jan 1, 2016

By I. J. Burgess, B. C. Barlow

"The electrodissolution of nickel pellets with different controlled electrical variables was investigated in an effort to mitigate plating residue in Watts bath based electroplating cells. During galvanostatic dissolution at 0.4 A dm-2, the anode potential was seen to oscillate between passive and pitting potentials. In-situ microscopy and ex-situ characterization has shown that lacy-pit formation and large, partially perforated, pit covers form during the potential decay from pitting potentials. In general, covered pitting at low current density shows increased plating residue. Using potential pulses is one way to maintain greater control over surface pitting by creating a higher density of pit nuclei while at the same time controlling their growth. Pits generated from a series potential pulses were examined ex situ for a range of pulse times and stages of pit growth were identified. An alternative method of anode dissolution was devised using potential pulses and duty cycles that allowed average current densities comparable to industrial plating conditions to be achieved. By restricting the time at pitting potentials to 0.05 s, plating residue was reduced more than six-fold in comparison to an equivalent galvanostatic experiment run at a comparable current density.INTRODUCTIONIn typical galvanostatic (direct current) nickel electroplating cells employing soluble anodes, some amount of plating residue is produced which consists of metallic nickel and impurities that detach from the bulk in the form of small fines. Significant buildup of residue requires plating operations be halted and the anode assembly cleaned of residue before operation can resume. While variables in the refining process have been shown to affect plating residue, more precise control of electrochemical variables may provide an attractive route to mitigation of residue."

Jan 1, 2016

By W. Andrew Falls

Although we have yet to reach the midway point in the decade, the copper concentrate market has already witnessed some of the most volatile swings in history, moving from record high treatment and refining charges in 1991-92 to historic lows, in real terms, during the past year. What has been the cause for such volatility? By and large the swings can be attributed to factors of supply and demand in the "custom" copper concentrate market. For example, two relatively recent events which had a major impact on the supply-demand balance were the shutdown of the Bougainville copper mine in Papua New Guinea in 1989 due to rebel insurgency and the startup of the La Escondida mine in Chile in 1992. The first event, threw a balanced to slightly surplus concentrate market into a substantial deficit, forcing treatment terms down as smelters, caught short by the loss of concentrate supplies scrambled to cover their positions. The second event, brought over 300,000 mt per year of new copper in concentrate to the market, more than reversing the losses from the Bougainville shutdown. In combination with other additions to mine capacity via new mines or expansions, the custom concentrate market moved into a huge oversupply in 1991-92 and many mines found themselves unable to place uncommitted tonnage. But even this event was quickly overcome by the use of underutilized capacity at non-traditional buyers, such as those in the C.I.S., Zambia, or those which were not traditionally buyers of custom materials. All of this has happened in the space of five short years, a mere snapshot in the history of the market. I'll return to these points a little further on, but first I would like to review the fundamentals which govern the copper concentrate market.

Jan 1, 1995

By Francisca San Martin, Tomas Vargas, Willy Kracht

"In Cu-Mo deposits it is usual to find considerable amounts of pyrite (FeS2). When processing this kind of ore by flotation, pyrite is rejected using lime to increase the pH to alkaline conditions (pH ~10.5). In the case of using seawater without pre-processing, i.e., raw seawater, the lime consumption increases dramatically and the recovery of molybdenite drops due to the precipitation of magnesium hydroxide on its surface. In order to avoid these negative effects, alternative ways of depressing pyrite should be considered.The current work introduces the use of Acidithiobacillus ferrooxidans, bacteria commonly used in bioleaching, as an alternative to depress pyrite in seawater flotation. The paper presents the results of biodepression in three systems: fresh water, water with added sodium chloride at 35 g/L (which corresponds to the salt concentration in seawater ) and seawater. The results show that it is possible to bio-depress pyrite with A. ferrooxidans in natural seawater pH (7.8-8.2). A similar behaviour is observed when running the experiments in fresh water, and in water with sodium chloride added.INTRODUCTIONWater is a scarce but essential resource for human development. The main reason for this scarcity is that, although 70% of the earth's surface is covered with water, only 3% of the total is fresh water and an important part of this water is frozen, so that only 1% of fresh surface water can be used for human activities. In Chile, most of the mining activity is concentrated in the north of the country, located in the Atacama Desert, considered the world's driest desert. That is why some mining companies have chosen to use seawater in their processes. For example, Esperanza (Antofagasta Minerals), and Sierra Gorda (KGHM International) use raw seawater in their flotation processes."

Jan 1, 2016

By Simon Pelletier, Jose Delgadillo, Dave Page, Pramod Kumar, Raj Rajamani, Dilek Alkac, Martin Fillion

"In general, many past works examine mill load, mill power, shell lifter design etc with a view to improving the mill capacity. On the contrary, only a handful of papers have addressed the transport of fine slurry through the pulp lifters. We present the flow dynamics of pulp lifters via laboratory scale experiments, high speed photography and modeling with computational fluid dynamics. Next, the flow behaviour in ArcelorMittal’s Mont-Wright autogenous mill discharge flow is examined. In an effort to improve mill throughput, a comprehensive simulation and plant scale study of the mill was taken up with a focus on the shell and pulp lifter design. The shell lifter was found to be at its optimum limit for breakage of ore particles given the constraints of wear life and occupied volume. Next, the pulp lifter study was taken up with computational fluid flow code and discrete element code. The simulations directly showed that there was considerable carry-over of slurry. This effect was traced to the area restriction in the pulp lifter channels. Further, crash-stop studies confirmed slurry pooling within the mill. Thus it was readily apparent that mill throughput can be increased if the pulp-lifter could be redesigned. Consequently, the discharger section of the pulp-lifter was redesigned.First, this paper outlines the modeling in laboratory scale studies and plant scale studies leading up to the discharger design. Next, the discharge flow in seven different published mill surveys is predicted with computational fluid dynamics (CFD) methodology shown here. Finally, Mont-Wright mill circuit performance before and after the discharger redesign is described. The usefulness of fluid flow simulations in the design of pulp-lifters is emphasized."

Jan 1, 2012

By G. A. Kosick, M. Freberg, L. A. Kuehn

"Highly variable and higher than forecast lead miff feed grades were problems experienced at the Polaris Concentrator. Limited flotation capacity and poor operating flexibility prevented the plant operators from overcoming these difficulties, and resulted in grades and recoveries which were lower than necessary.Two column cells were integrated into the lead circuit in an effort to give the operators the tools they required to achieve the best grades and recoveries possible, regardless of the miff feed grade. Installation of the column cells increased lead concentrate grade from 76% to 79.3% Pb, and increased lead recovery from 9/% to 91.3% Pb.This paper reviews the design, development and operation of column flotation cells in the lead circuit at the Polaris Mine.IntroductionThe Polaris Concentrator was constructed in 1980 on a barge in Trois Rivieres, Quebec, and towed to its pre sent location on Little Cornwallis Island, Northwest Territories, about 1450 km from the North Pole. The plant presently treat s about 2800 mtpd of new feed with average feed grades of 3010 Pb, 13% Zn and 5% Fe. Current production averages 36 000 tonnes of lead concentrate and 220 000 tonnes of zinc concentrate annually.The barge design and min e location dictated that the concentrator .be compact with very little unused floor space. As a result, flexibility was limited in terms of expanding the plant.Mill feed grades were highly variable and above forecast and the lead flotation circuit was found to be under capacity. This caused continued overloads in the lead cleaning circuit, resulting in low concentrate grades and recoveries. It was evident that an increase in lead flotation capacity and flexibility was necessary. The existing floor space did not allow for the installation of additional conventional flotation cells, but one alternative was the installation of column cells."

Jan 1, 1988

By F. Duan, J. Pfarr, J. M. Goris

"The U.S. Bureau of Mines, in cooperation with the Homestake Mining Co. and the University of Utah, is conducting a field demonstration to evaluate the behaviour of rock masses supported with cable bolts. One of two rock masses being studied is supported with both conventional and birdcage cables. This rock mass is located between the 6500 and 6650 levels of the 45-48N stope, 21 ledge. The second rock mass is the hanging wall of the D-limb stope between the 3500 and 3650 levels near the Ross Shaft and is supported with conventional cables. The primary objectives of the project are to monitor the behaviour of the two rock masses during mining using extensometers and cable support strain gauges, after which the rock will be modelled using the finite-element code UTAH III. Field data will be used to verify the modelling. In addition, a comparison between the effectiveness of conventional and birdcage cables will be made.45-48N StopeThe 45-48N stope is located in the 21 ledge between the 6500 and 6650 levels and contains about 233 000 short tons of mineable ore. The orebody lies within the Homestake Formation and is very irregular being approximately 200 ft long and 75 ft wide. Between the 6500 and 6650 levels, it dips from 35 degrees to 60 degrees east with a near-vertical plunge to the south. Mineralization is characterized by arsenopyrite and pyrrhotite with minor gold. The contact between the Homestake and the overlying Ellison Formation creates a plane of weakness along localized graphitic layers. The footwall dips at approximately 60 degrees but between 100 ft and 120 ft above the 6650 level , the hanging wall flattens to almost 20 degrees causing a significant mining problem."

Jan 1, 1991

By Jessica Hurst, Mark Holdsworth, Jeff McKay

"In 2002, Kennecott Utah Copper Copperton Concentrator installed an Expert Control system for optimization of their SAG and Ball milling circuits. The advanced process control has been well received and is utilized well over 95% of the time. In the years since, many lessons have been learned and many advances have been made. Improvements have produced millions of dollars of benefits and have included directly controlling feedrate with complex logic as opposed to using a bearing pressure controller, modifications to how sound is used for control, modifying logic for maximizing asset utilization and using advanced control for automating equipment restart. INTRODUCTIONRio Tinto’s Kennecott Copperton Concentrator produces copper, molybdenum, gold and silver. The mill has 4 SAG mills (SAG 1-3, 34 foot diameter and SAG 4 36 foot diameter), 8 ball mills and copper and molybdenum flotation plants, as seen in Figure 1. The 2010 average data is shown in Table 1.In 2002, a Metso Cisa OCS Expert Control system was installed at Copperton for SAG and Ball mill advanced control. The expert control was first placed on SAG 2; data from the first few weeks compared with operator performance indicated approximately an 11 percent increase in throughput. After full commissioning, the system was gradually refined further by plant personnel to develop key performance indices based on SAG feed size, refinement of the sound setpoint for control of mill speed and throughput, specific rules to help the adjustment of speed based on loss of a feeder, etc. With the addition of the pebble crushing circuit, the expert control began utilizing a model estimation of crushed pebble bin level to help stabilize the impact of returned crushed and uncrushed pebbles on the SAG mills. As the ore and limitations change, it is necessary to evolve the SAG logic to better handle any constraints."

Jan 1, 2012

By Geo. D. Furse

"The information that has accumulated regarding the geology of the McIntyre mine is the result of work by several geologists and, as might be anticipated, their conclusions have varied. The writer has been at the mine only a few years and due to wartime conditions has not been able to study and review the mine's geology as thoroughly as his predecessors. Where the writer's conclusions differ from those of previous authors, it should be noted that he has been unable to re-map all the workings in which flow contacts are exposed, but has endeavoured to describe the structure as indicated by the evidence he has been able to observe. He wishes to thank H. G. Skavlem and H. S. Robinson for their constructive criticism of this paper, and McIntyre Porcupine Mines, Limited, for permission to present it. A list of references is appended.The McIntyre mine is in the central part of the Pearl Lake section of the Porcupine gold area. It includes the ""bed of Pearl lake, twenty-two claims to the north of the lake, and the Platt Vet lot to the southeast. It is bordered on the west by the Hollinger and on the east by the Coniaurum property. The chief ore-bearing veins on the upper levels lie very close to the north shore of the lake, but some occur under the lake, and others of lesser importance are found to the southeast. The most important veins are Nos. 3, 5, 7, 10, and 25. Some of the veins extend into the Hollinger. on the west and others cross the east boundary into the Coniaurum. The total production since the commencement of milling in 1912 to the end of March 1947 has been $168,448,218 in gold and silver from the milling of 17,690,963 tons of ore."

Jan 1, 1954

By W. A. Crosby, W. K. Smith, K. L. McDonald

"IntroductionThis paper describes how the introduction of two blasting techniques and the resuItant development of new dragline digging methods have increased the productivity of single- and double-seam dragline operations at the Rietspruit opencast mine.LocationRietspruit , one of South Africa's largest opencast strip mines, is situated some 11 5 kilometres east of Johannesburg in the Witbank coal field.BusinessRietspruit is a joint venture between Transvaal Consolidate d Land and Exploration Company (T.C.L.) and Shell, South Africa. Rietspruit is administered and managed by Rand Mines, the Mining division of Barlow Rand, transportation and marketing being the responsibility of Shell. The total saleable coal product, approximately 5,5 million tons per annum, is exported via Richards Bay.Presplitting and Throw BlastingPresplitting was initially introduced in September 1980, as a means of dewatering t he overburden and mid burden to facilitate the use of dry-mix explosives a s opposed to slurry explosives.The utilization of the throw blasting technique arose as a direct result of observed ground movement after the successful introduction of presplitting. The first trials commenced in March 1981 on the double-seam operation. The methods, economics and resulting advantages are fully described elsewhere. A brief description follows.Method of PresplittingAfter determining the desired presplit line, 250-mm holes are drilled to top of coal at 4-m to 6-m intervals, dependent on the competency of the rock overlying the coal. A single 60-kg charge of slurry explosives is located at the bottom of each hole, no tamping is used and all the holes are detonated simultaneously. The results are a competent, straight, vertical highwall. This enables accurate control of the first row of blast holes closest to the highwall, which is a prime requisite for successful throw blasting."

Jan 1, 1983

By George Dogbe

Open pit optimization forms the basis of mine design, production planning and construction of mine facilities. The negative effects of this phase may include substantial reductions in ore recoveries, relocation of waste dumps and surface infrastructure, extraction of unwanted waste, with a cumulative effect on mine value reductions. As a result, this area has received significant research focus using linear and dynamic programming, heuristic, genetic and stimulation algorithms and the graph theory. These algorithms are derived from rigorous mathematical models, but they are built on the assumption that all the periodic production capacities are extracted at the time of analysis. This assumption weakens the models because mining occurs over significant time period. In order to develop a comprehensive model that accounts for the timing of material extraction, the current models, which are built on the economic block value (EBV) concept, must be modified to account for time. In this paper, the authors develop a model of the optimization problem using the 2-D Lerchs-Grossmann?s optimization algorithm. This model is used to investigate the effect of periodic discounting on the optimal pit layout design by designing an integrated pit optimiser (IPO) that incorporates ?production scheduling?. IPO uses the ensuing discounted economic block value (DEBV) as a basis for the pit optimisation process. The time-dependent model is compared with a time-independent model. The results show that discounting has no effect on the optimal pit contour although the reported pit value is overstated if the EBVs are used instead of the DEBV. The authors, however, believe that this conclusion may not be general and could have been influenced by the nature of ore body geometry and the proposed mining strategy.

May 1, 2004

By Graham Swan

The case is presented for the development of a spray-on liner for ground support, but only in the context of a new mining system, consistent with the technical and business case advantages of liners compared to alternatives. Over the past 10 years much effort has been expended in specifying and testing the mechanical/environmental properties of various liners, without actually fulfilling any mining system requirements. The results have been generally disappointing, with no consistent and sustained application in a dayto- day production environment being achieved, anywhere. Central to this shortcoming has been the failure of mining/liner ma nufacturing companies to realise that screen/mesh or shotcrete cannot simply be replaced with thin, deformable liners: success will depend upon closely defining and evolving a new, liner- functional mining system. The paper begins with a discussion of some fundamental technical capabilities and constraints associated with 3M Canada?s liner material that has been developed in partnership with Falconbridge Limited. Intended applications for this liner include the support of bursting ground associated with deep mining conditions. With safe application and consistent performance in mind, the paper argues for a new mining system that derives in the first place from best quality drilling and blasting practices. Taking this as a given, a system for waterscaling and liner application prior to mucking out is proposed, followed by bolting through the liner while arranging for the bolts to be explicitly connected to the liner. Some logistical and equipment related challenges of such a mining system are then reviewed and discussed in the context of the business case for liners. It is concluded that this approach, if successful, will provide the greatest benefits in the areas of safer mining, due to reduction in worker exposure, and in reduced mining cycle times.

May 1, 2003

By J. G. Kim, B. G. Chae, D. S. Cheon, E. S. Park

The underground structures such as high-level radioactive waste disposal thermal energy storage should be designed and constructed considering thermal-mechanical characterization because of long term and/or cyclic thermal stress. High temperature changes in these structures can extend the existing cracks and generate new cracks in surrounding rock mass. In case of long term and/or cyclic thermal stress conditions, it is possible to lose the function of structures caused by the significant cracks and fracture growth in the rock mass. In this study, the critical thermal crack temperature of the granite, limestone and thermal Kaiser effect were evaluated. Experiments were carried out with the heating speed of 1.5?/min. To avoid thermal shock and target temperatures were set at 150? and 250?. In order to verify the thermal Kaiser effect, the samples were heated by 150 ? after 2-3 repetition to 250?. The generation and growth of cracks were observed through an acoustic emission measuring system. The variation of the physical rock properties was measured before and after thermal experiments. After compared with the thermal experiments using PFC2D, numerical simulations were carried out at higher temperatures (400?). The experiments showed that the initial thermal crack temperature of granite and limestone is about 35?, 32? and the critical thermal crack temperature were about 69? , and 49? , respectively. The experiments confirmed the thermal Kaiser effect of experimented granite and limestone due to thermal cyclic loading. From the numerical analysis, the critical crack temperature showed the same tendency as the thermal experiments. The P and S wave velocity of the sample after the experiment were reduced. Therefore, when heat is applied in long-term and repetitive conditions, these cracks may affect the function and stability of the underground structure. Therefore, further studies of thermal stress effects should be performed.

Jan 1, 2015

By SangHyeok Han

Selecting the best possible cranes and identifying spatial conflict-free locations on sites can result in productivity and safety improvements for large-scale industrial construction projects. In the current practice, experienced lift engineers select cranes based on the heaviest lift and/or the largest lifting radius of the identified crane. This practice is relatively time-consuming, and optimization of the crane?s use and location is also difficult. There are many factors which need to be considered during the crane selection process, a reality which further complicates the process. This paper presents a framework which aims at developing a decision support system to enhance the crane selection process and collision-free path planning for large-scale construction projects. This paper utilizes an innovative crane selection matrix in order to establish a process for optimized crane selection for construction projects. The study considers more than 40 different factors in order to reduce time and improve safety for crane operations. Following finalization of crane type (mobile crane versus tower crane), a visualization model to simulate crane operation and identify collision-free crane operation paths is proposed. This process can assist project managers to plan the lifting process more effectively and efficiently. A case study-based approach was utilized to demonstrate the proposed methodology. The methodology was tested in the planning and construction process for boiler house structures in Mannheim, Germany. The project entailed numerous challenges: one of the major tasks was to lift a 102-ton load on the top of the boiler structure through crane collaboration; space limitations on site also presented several challenges related to crane selection, location, and operation processes. Based on the project constraints, the proposed crane selection framework, and the visualization models, two tower cranes were selected and successfully implemented in the case study.

Aug 1, 2013

By Tim Gartner

This paper introduces a new application of a proven technology to provide dynamic, voltage regulation in underground mine electrical distribution systems. This technology acts to ?stiffen? the electrical power feeding continuous miners and generally raise and improve the voltage regulation at the continuous mining machines which may be located over long cable distances from the mine incoming electrical substation. This is extremely valuable for large underground continuous mining operations that use high power electrical mining machines fed over very long cable lengths (measured in 10s of km, from mine shaft to mine operating face). This new technology provides fast acting dynamic voltage regulation (voltage support) while starting continuous borers and extensible conveyor belts. Additionally, if the mine suffers from a lower distribution voltage due to long cable lengths, this new technology will raise the overall voltage at the mining machine, which can increased mining machine production and lead to longer motor life. The technology is based on voltage source converters operating as static VAR compensators specially packaged for the harsh underground mine environment. While static VAR compensator technology has been successfully used in utility and large industrial applications for many years (including open pit mining and arc furnaces), it has not (until now) been applied in underground mining. The paper will present the problem of poor voltage regulation in underground mines on weak electrical utility networks. It will then discuss some of the alternatives that have been used to resolve this voltage problem. It will then introduce the new technology used to solve the problem and review the issues involved in using this technology in the harsh underground environment. Finally, the paper will present preliminary performance results from an actual installation in a large underground continuous mining operation in Canada.

May 1, 2003

By Stephen G. Hardcastle

Worldwide, metal mines are going deeper. For example in Canada there are six such mechanized mines planning production at 3000m (10,000?). Working at such depths challenges all aspects of mining including the provision of ventilation to supply an equitable working climate for personnel and machinery. This paper explores the overall challenges of supplying ventilation in Canada?s deepest mines and how the volumes required could be minimized. This is essential because it must be remembered that ventilation and refrigeration are not only costly items to supply, but the magnitude of their cost is also very sensitive to the volume supplied. For ventilation this is obvious because of the cubic relationship between volume and cost. In deep mines, the main function of ventilation is the removal of heat transferred from the strata and generated by the mining machinery, and the air volumes supplied are designed accordingly. The second consideration is the removal of other pollutants that could be harmful to humans such as diesel and explosive fumes. However, the need for ventilation at its present design levels could be reduced. This paper discusses the implications of such technologies as: teleremote mining and alternatives to diesel equipment on the magnitude of the volume supplied; ventilation-on-demand which recognizes ventilation as a costly and limited resource; plus other aspects, such as the benefits of controlled recirculation, local cooling, and heat rejection back to the strata to gain further advantage from new developments. Furthermore, this paper advances the need to be able to design ventilation systems more cost effectively and how such new advances as process simulators can be used in conjunction with ventilation simulators to explore the diverse options that could affect the air volumes required and provide more realistic economic data upon which to base design decisions.

May 1, 2003

By J. C. Salas, A. Weishaupt, J. A. Guzman

The origins of the electrorefining process date back to 1869, when the first ER cathodes were produced at Pembrey, South Wales. By the time that Pembrey was closing, Chuquicamata tank house was under construction, initiating in 1915 a long history of electrolytic cathodes production. Initially electrowon cathodes and later electrorefined cathodes were produced. During the history of Chuquicamata tank house, there have been several challenges associated with the quality demanded for electrolytic copper. Among them, the transformation from electrowinning to electrorefining in 1952, the onset of the direct commercialization of ER cathodes and the phase out of wire bars in the 80s, several technological changes including the implementation of permanent cathodes, and certainly the constant increase in complexity of the anodes refined. This paper explores the manner in which Chuquicamata has addressed these challenges and examines its links to process characteristics, which have been necessary to maintain a product quality according to market requirements and a brand, “CCC”, that is widely recognized worldwide. INTRODUCTION Located in the north of Chile, in the Atacama Desert, Chuquicamata is an open pit copper mine that for many years has played a relevant role in the global and local mining industry. Known from prehispanic times, the Chuquicamata deposit was initially developed between 1912 and 1915 by the Chile Exploration Company, the operating company of the Chile Copper Company controlled by the Guggenheim Brothers in the USA (Yeatman, 1916). Officially inaugurated on May 18th, 1915, by the President of Chile who pressed a button in Santiago sending a telegraph signal to the power generation plant specially built in Tocopilla, 1500 km north. It signaled the beginning of generating energy, sending it to Chuquicamata, another 140 km east (“El mineral de Chuquicamata”, 1915), marking the beginning of its industrial operation.

Jan 1, 2019

By A. Haque

"This paper describes an improved formulation of an analytical shear strength model developed by Haque & Kodikara (2012) by incorporating the failure of regular concrete/rock joints subjected to CNS condition. Sokolovskii’s analytical solutions developed for failure stress of slopes have been found to be effective in this regard. The analytical models have been validated for predicting the shear behaviour of concrete/rock joints subjected to various CNS loading conditions. The model has been found to accurately predict the dilation, normal stress and shear stress behaviour for a regular triangular saw-tooth joint up to the peak shear stress. Moreover, the peak shear stresses of concrete-rock joints with asperity angles ranging from 5 to 27.5o compared reasonably well with the available laboratory test results.INTRODUCTIONA new shear strength model for elastic regular triangular concrete/rock joints loaded under Constant Normal Stiffness (CNS) boundary condition has been developed by Haque & Kodikara (2012). The salient feature of this model is its capability of accurately estimating the normal and shear stresses of joints with shear displacements under CNS condition by incorporating the elastic deformation of asperities. However, the shear strength model of Haque & Kodikara (2012) for the elastic asperities cannot be directly employed to predict the peak shear stress of a joint experiencing plasticity under applied loads. Therefore, the model inevitability requires articulation of an appropriate shear failure criterion simulating the boundary conditions of the loaded asperity to predict the peak shear stress and corresponding normal stress and dilation behaviour of concrete/rock joints. Sokolovskii (1960) developed a set of analytical solutions for failure stress of slopes comprising c-??soils, which can be exploited effectively to simulate the failure condition as encountered for loaded regular triangular concrete/rock joints (Seidel & Haberfield, 2002)."

Jan 1, 2015

By F. Q. Barnes

"Uranium, thorium and yttrium have been recovered, since the 1950s, from pyritic quartz-pebble conglomerates occurring in the basal sediments of the Quirke syncline at Elliot Lake. Gold, an element of interest in similar type deposits occurring worldwide, is a very minor constituent of the Elliot Lake conglomerates because the provenance area of the sediments had few gold-bearing rocks.Uranium, thorium and rare elements are found in resistate and diagenetic minerals occurring in the matrix of the conglomerates. The valuable elements are recovered from the acid soluble minerals during the milling process.Large uranium reserves occur in mines found on both limbs of the Quirke syncline. Major reserves are being developed on the South Limb of the syncline, where the Stanleigh Mine, among four past-producers, has been reactivated.IntroductionUranium and thorium commonly occur in anomalous amounts in coarse basal sediments. Radioactivity in basal conglomeratic sediments is generally attributable to resistate minerals, that rarely occur in concentrations high enough to be of economic significance. One type of basal conglomerate found worldwide, in Precambrian cratonic basins, has uranium or uranium and gold concentrations large enough to be of economic interest. Principal among the occurrences are the pyritic quartz-pebble conglomerates of the Huronian in Canada, of the Witwatersrand in South Africa and at Jacobina in Brazil.The uranium and gold mineralization of these conglomerates was derived from provenance areas where both elements occurred and were later concentrated through source degradation and deposition in placers in a reducing environment. Only the Huronian conglomerates, generally much richer in uranium but with negligible gold, have been exploited solely for their uranium content.The uraniferous conglomerates of the Huronian at Elliot Lake, 500 km north of Toronto, continue to represent a major world source and supply of uranium concentrate."

Jan 1, 1986