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By Timothy D. Dreger

In the spring of 2009, Suncor Energy Inc. technical personnel saw an opportunity to improve planning 0and mining of overburden material in order to address aggressive construction goals and increased mining complexity. By tying together current and historical equipment performance, geological model information, and material observation in the field new operating elevations for the shovels were developed. At the same time, methodology of recording and reporting overburden movement through the Modular Mining DISPATCH system was changed. The result of this first stage of improvements was an increase in construction-grade material available at the face. The personnel involved realized that other opportunities were available to build on the initial successes and establish processes that would stand up to attrition and experience levels. A cross-functional technical team was assembled to identify ways of further improving workflows around overburden mining. Some of the opportunities realized include greater alignment between planning, reporting and operations as well as development of reporting tools that help steward Mine Operations to the plan. All of which have resulted in Suncor meeting or beating its forecast overburden targets. The cross-functional technical team continues to seek out new and better ways of doing business so Suncor can continue to deliver on its commitment of outstanding performance and results.

May 1, 2011

By Masafumi Maeda

Metal recycling and waste treatment have become common issues in various metal industries, notably, the iron and steelmaking industry and the nonferrous industry. The former tends to treat mass wastes such as municipal wastes or plastics. However, volatile compounds are condensed in dusts of their incineration processes and they should also be treated, likely by the nonferrous industry. Those "end-of-life" products contain heavy metals that might be harmful to mankind and should be handled by the non-ferrous industry. Not only revenue derived from regeneration of metals from the secondary materials but also the revenue generated from the treatment charges is a strong attraction for nonferrous industry to process secondary materials. The present study presents some statistical data of the Japanese corporations in the iron and steelmaking industry as well as the nonferrous industry on their revenues generated from processing of secondary materials. Some of the companies are now concentrating their business resources on new areas of the market. Although the quantity of waste or recyclable material is still small at present, the quantity will increase in the near future, and it is anticipated that international collaboration is required to handle the greater variety and the larger quantity of secondary materials.

Jan 1, 2004

By M. Moghadami

Design and construction of underground openings are associated with many uncertainties with regard to input data. In the construction process of underground openings various levels of risks exist and these affect the safety and profitability of the project. The development of underground openings can be broadly categorized into three major phases of data collection, design, and construction. A significant level of uncertainty is associated with some of the key parameters involved in the aforementioned development phases. Moreover, it is very difficult to evaluate and quantify the effects of uncertainties on the underground project construction process. Additionally, the interactions between important risk elements are typically not well understood. The intention of this study is to develop a general risk model for underground opening projects located in soils and weak rock masses. The goal of the research is to identify and quantify important geotechnical risk elements. Moreover, a weighting procedure will be proposed for the identified risk elements. Finally, all important risk elements will be assembled into a general model with emphasis on geotechnical parameters. The final objective of this study is to apply the proposed model to a real underground project scenario and verify it for actual field conditions. The proposed model, once calibrated properly, will provide a useful support tool for the decision making process.

Aug 1, 2013

By R. H. McCorkell

"Sample drops were evaporated on Lexan disks on which fission tracks were etched and counted after neutron irradiation. [U] = 0.01 ppb could be measured. This method is rapid, cheap, simple and accurate when sufficient control and checking is possible, but several types of error were encountered in its use, the natures and sources of which were investigated but not always determined. The most troublesome were ""annealing"" processes which sometimes altered the fission tracks so that they were not fully revealed, or acquired varying appearances, after etching. Several different annealing processes seemed to exist and seemed to be produced by alteration of the Lexan surface by dissolved components in the solutions and, perhaps, by radiation and local heating during neutron irradiation. The appearances of the etched tracks varied with the natures of the droplet solutions even when annealing appeared to be absent, and for this reason, standards are best prepared by additions of known amounts of U to sample solutions. Contamination of sample droplets by U-containing dust from the air is best avoided by preparing the disks away from areas where other U-analysis is being carried out.IntroductionFission tracks are regions of radiation dam age produced in solids by the fission of atoms such as 235U. They can be enlarged to sizes visible through optical microscopes by etching surfaces through which they pass with solvents in which the solids are slightly soluble. A sample to be analyzed for U may be placed against a film of a solid of negligible uranium concentration (the track detector) and irradiated with thermal neutrons in a nuclear reactor to fission 235U. The track detector is then etched and the number of tracks per unit area, ?, which is proportional to the uranium concentration of the sample, c, and the neutron do se, n, is found by counting through a microscope. Uranium in solutions ha s been measured by evaporating drops on track detectors, irradiating, etching and determining the number of tracks per drop, T(2,3,4,5). Extensive test s of this method are reported here."

Jan 1, 1983

By David A. Fawcett

"This paper presents a summary compilation of current Canadian coal research projects in the areas of exploration, mining and preparation. The compilation forms the basis for a review of present operations and technology, and of current research efforts. General research areas, requirements and priorities are suggested along with a discussion on achieving the research goals in terms of funds, facilities and staff. The current level of research funding in these areas is assessed. IntroductionCanada's coal production is expected to undergo major expansion over the next two decades in response to domestic demands on coal for electrical power generation, industrial uses and conversion to synthetic .fuels, and in response to export demands for both thermal and metallurgical coal s. Surface mining will continue to dominate, but underground mining will contribute an increasing proportion to total production.In Canada, coal research, particularly in the areas of exploration, mining and preparation (E, M & P), has been limited. Increased impetus has recently been given to this research with expansion of CANMET's activities and with the establishment of the Coal Mining Research Centre (CMRC). University researchers are branching into coal research; strata mechanics and preparation are popular in mining-related fields.With the increase in number of organizations and individuals involved in coal research, a means of keeping abreast of the ongoing work is necessary. An annual summary of the status of and developments in current research assists in providing this information. The CMRC prepared such a summary for Canadian coal E, M & P research in 1979 and it received favourable response, justifying the effort. A 1980compilation has been prepared and pro vides the basis for this paper; it is presented with a brief discussion on background, format and content.The compilation forms a basis for an examination of the current status of Canadian technology in E, M & P and of the ongoing research in each area. General research requirements are reviewed and programs and tasks are suggested, followed by a discussion of research goals in term s of priorities and achievement. An assessment is made of the level of research funding in Canada, with comparisons drawn from the United States, the United Kingdom and Australia."

Jan 1, 1980

By Wataru Inagaki, Kozo Kawano, Tetsuya Ando

Cu–Ni–P alloys are the typical precipitation-hardening material. Thus, Cu–Ni–P alloys is expected to apply alternative materials for heat-exchanger tubes. Therefore, it was investigated that the effect of the heat-treatment simulated brazing on mechanical properties of Cu–Ni–P alloy in this study. The hardness of specimen as heat-treatment simulated brazing after solution heat treated (sht) was shown approximately 105 HV. The peak hardness and tensile strength of the specimen as performed aged at 498 K for 43.2 ks following the heat-treatment simulated brazing achieved at approximately 140 HV and 260 MPa, respectively. From these results, it is expected the application to higher-strength tube for heat-exchanger after brazing in a furnace. INTRODUCTION Copper prices have increased dramatically with high in recent years. At the same time, manufacturing industries are required to reduce the environment load by reducing the wall thickness of the copper tubes used in heat exchangers. High-strength copper alloys are developed for reducing the wall thickness, while achieving high strength, of the copper tubes. Moreover, excellent workability and corrosive durability are also required by these copper tubes. Furthermore, the brazing, which are carried out at about 1050 K during the assembly process of heat exchangers, leads to grain growth and re-solving of additive elements. Nomura et al. achieved a tensile strength and Vickers hardness of approximately 260 MPa and 110 HV, respectively, after aging at 723 K for approximately 36 ks. Tamagawa et al. studied that the strength of Cu–Ni–P alloy rapidly increased at about 125 HV during the cooling process after the heat brazing due to the Ni2P precipitates sized 5–10 nm. Furthermore, Inagaki et al. achieved a Vickers hardness of approximately 130 HV by aging at 498 K for 10.8 ks. In the present study, we focused on the lowest aging temperature of 498 K in the previous study, and investigating the effect of heat-treatment simulated brazing on the mechanical properties of the Cu–0.93 mass%Ni–0.24 mass%P alloy.

Jan 1, 2019

By M. H. Moys, C. Bhondayi, Eric Tshibwabwa

"The relationship between froth bubble size as represented by intra-bubble impact distances (IID) and flotation performance is investigated. Experiments using an artificial ore comprising of limestone as the floatable component and silica as the gangue material were carried out in pseudo-steady-state circuit. A method to measure IIDs developed by Bhondayi and Moys (2014) was then used to measure froth bubble sizes as a function of froth height, superficial gas velocity and feed rate (pulp residence time). Samples for flotation performance evaluation were also taken. Results indicate that as froth height increases, the average IID values increase resulting in a decrease in limestone recovery while the grade increases. Increasing gas rate results in lower IID values that are accompanied by high recovery and low limestone grade. The effect of feed rate on both flotation performance and average IID values depended on froth depth and gas rate. Average IIDs largely remained the same at the two feed rates investigated.INTRODUCTIONIt is regarded as obvious that froth bubble size distribution (FBSD) will have an effect on plant grade and recovery. As the average FBSD decreases, the bubble surface area flux will increase making possible an increase in recovery and a decrease in grade. The opposite trend will be involved if average FBSD is increased. However: no method for measuring FBSD has been available, so these trends have not been studied in any detail. Bhondayi and Moys (2014) developed a novel technique to measure or estimate froth bubble sizes. Although they established a potential to use this method in understanding froth performance, no data have been presented to assess the correlation between the estimates of bubble sizes and actual flotation performance. In this work experiments are done to establish how the change in froth phase bubble sizes affect flotation performance. Flotation performance will be evaluated using recovery and grade. How water recovery to the concentrate changes with changes in bubble sizes will also be evaluated. It is known that water recovery is approximately inversely proportional to the square of the overflowing bubbles sizes. It will be interesting to establish how the rate of change of bubble sizes changes with water recovery."

Jan 1, 2016

By P. O. Rag, Daniel E. Walsh

"Gravity concentrators are evaluated either in closed or open circuit. The process developed by MIRL employs the use of neutron activated gold (/98Au) and radiation detection circuits which monitor the concentrator's product pulp streams. Concentrator recovery is determined from the ratio of the number of gold particles detected as passing into the concentrate to the total number of gold particles detected over all monitored pulp streams. This system eliminates the experimental errors introduced by sampling, sample reduction, and assaying and allows recovery to be evaluated for gold particles of specific sizes, shapes, and flatness. In addition to allowing for a more accurate and detailed recovery analysis, this process significantly reduces the time and cost required for unit evaluation when compared to other gold recovery analysis procedures. This system enjoys its greatest utility where a large number of tests are to be run in order to evaluate a concentrator over a variety of operational settings, i.e. factorial designs, and allows accurate testing of concentrators operated in closed circuit systems.IntroductionThe surge in gold prices of the late 1970s and early 1980s was followed closely by a renewed interest in gravity concentration processes, especially in concentrators claiming efficient recovery of fine placer gold. Cones, spirals, spinning bowls, jigs and cyclones are all considered potential fine gold concentrators. Fine gold, in this context, is defined as not only that gold finer than 250 um (60 mesh""), but also coarser gold, whose low shape factor may decrease its recovery by conventional, large-scale sluicing operations.Adoption of fine gold concentrators by the Alaskan placer milling industry was retarded by the lack of reliable, accurate, and reproducible data on concentrator efficiency with respect to fine placer gold recovery. In most cases, concentrator performance was based on a limited number of tests run on placer material, about which only slight knowledge of gold size and shape characteristics existed. In order to be prepared for the resurgence of interest in fine gold concentrators, which will follow the next upturn in gold prices, a detailed study of the fine gold recovery characteristics of gravity concentrators needs to be conducted."

Jan 1, 1986

By G. Kachaniwsky, M. Bédard

"The Noranda Continuous Smelting Process at the Home smelter was commissioned in March 1973. The ability of the process to handle a wide variety of feed materials and a steady stream of technical innovations in recent decades has served to maintain the Home as an efficient and versatile custom smelter. This paper reviews recent technological development and capabilities in areas such as sampling and preparation of secondary and complex materials, as well as the smelting and related environmental control issues. Future trends and improvements are also addressed.IntroductionThe Noranda Home smelter commenced operations in 1927. It was originally designed to treat 910 tonnes per day of copper concentrates and ores coming from the new Home mine site. In the 19305, the company opened its doors to other mines in northwestern Quebec. Treatment of feed materials on a custom basis has, therefore, always been an important factor in the operation of the Home smelter. Throughout the years, the Home gradually increased its custom smelting capacity, while maintaining its production from the Home mine. By the early 1950s, 65% of the materials processed was from outside customers, and by 1974, this figure had increased to more than 85% . With the closure of the Home mine in 1976, the company extended its smelting operations beyond ore and concentrate, and began to treat more scrap metals and residues rich in copper and precious metal content.Corresponding with the ever-changing pattern of feed sources, the Home smelter has adopted technological improvements to meet new and challenging business conditions. The technology of the original smelter of 1927has been completely replaced over the years in the quest for improvement. One of these major changes occurred in 1973 with the start-up of the Noranda Process Reactor. This extremely efficient smelting system has been exploited to treat a wide variety of copper and precious metal (PM) containing custom materials. This paper reviews recent developments and improvements achieved at the Home smelter, in the receiving, sampling and treatments of copper and PM scrap and secondaries, with particular emphasis on the development of new technology and equipment in complex feed materials treatment."

Jan 1, 1991

By L. J. Surges, C. Bazin, D. Hodouin

"A kinetic model was developed for the conditioning and flotation processes of an industrial lead concentrate dezincing circuit. The models were incorporated into a flowsheet simulator. Simulation results demonstrate the advantages of split conditioning of coarse and fine particles and the potential of this approach for the production of the heuristics required for the development of intelligent control strategies.The conditioning model relates the proportions of activated and hydrophobic surfaces and residual reagent concentrations to activator and collector reagent flowrates. The model accounts for the solids and water flowrates, and for the solids grade and size distribution. The flotation model uses the predicted proportion of hydrophobic mineral surface, the aeration rate and the froth depth to estimate cell, bank and circuit performances. IntroductionFlotation has been the object of modelling and simulation studies for more than two decades. The process is usually modelled as a kinetic one in which the particle flotation rate constants are key parameters. Rate constants are related to the operating variables (pH, reagent flowrates, aeration rate, etc), however these relationships are seldom established in the literature.In spite of the large amount of literature on flotation modelling and simulation, few papers deal with the interaction between flotation rate constants and operating variables. This important gap, which was identified in the early seventies by Woodburn (l970), still exists. One of the first attempts to include operating conditions in a flotation simulator was made by King (l973), who explicitly introduced the aeration rate into an air bubble loading model. Lynch et al. (1981) used an empirical model to relate the proportion of slow floating particles to the ratio of collector per ton of ore. Such empirical relationships are difficult to calibrate and cumbersome to use when an activator and a collector are added to slurries in which particle size distribution and surface characteristics are changing. Eascur and Herbst (l982) relate in a fundamental way the mineral flotation rate constants to the aeration rate, froth depth, agitation and frother concentration, however the model does not account for reagent flowrates. Niemi et al. (1982) have proposed a model for the conditioning of minerals and included it into a conventional flotation simulator."

Jan 1, 1988

By Donald C. Craighead

"Overland conveyor systems to transport bulk materials from mines to processing plants or load-out stations have to be seriously considered as the viable alternative to truck and train haulage systems. With the ever-increasing price of diesel fuel, conveyors with their electrical drives generally prove to be the least costly when operating and maintenance expenses are analyzed for the life of the installation.IntroductionOverland conveyor systems are now considered as one of the viable alternatives which should be closely investigated when a mining company and its engineers initiate studies on methods of transporting raw materials from a mine to a processing plant or rail load-out station.This paper will generalize the overland conveyor and compare it with alternative handling systems. Each installation must be carefully studied to determine the best route, type of foundations river crossings; amount of excavation work for cut and fill, maintenance roads, etc. No two locations are the same nor are the principles of operation of an overland conveyor the same as those of a normal plant or mine belt conveyor. The proper time and effort must be put into the study of the system to evaluate the costs for comparison with other systems.The length of the overland conveyor depends on the layout of the mine and plant and the topography of the terrain between the terminals. Individual lengths may be up to 20 km, and a series of conveyors may extend 100 km or more.Advantages of Overland ConveyorsThe main advantage of overland conveyors is the low operating and maintenance costs which will be realized over its years of operation. Although truck haulage probably will permit lower capital investment during the initial phases of the mine development when the annual capacities have not developed to the design level, their operating and maintenance costs will be considerably higher. When evaluating the total costs per ton of material to be handled during the expected life of the operation and the projected increases for diesel fuel versus electricity, overland conveyor systems will generally prove to be economically more viable."

Jan 1, 1982

By Wojciech Naworyta

The development of the Gubin lignite deposit, located in western Poland is the first mining investment of this magnitude in 30 years. Gubin is the second largest prospective lignite deposit in Poland, with over one billion tonnes of indicated reserves, which would be sufficient for a 2700 MW power plant for almost 50 years. The paper presents select problems related to the mining development of Gubin deposit. Attention has been paid to changes in legal and formal requirements for mining investments that were introduced after the change of Polish social and economic system. The significance of spatial, environmental and social conditionings has been emphasized. Typical attributes of the mine planning process have also been presented. Those include the long time horizon of the investment, uncertainty of external conditions and resulting difficulties in the project?s economic evaluation. Focus has been placed on mutual connections between particular elements of the project and time dependence during the process of their development. Iterative characteristics of the design process have been proven, together with resulting necessity of frequent verification of previous assumptions. As a background for presenting selected problems related to future development of Gubin lignite deposit, the methods of design have been discussed ? the variants method, the empirical method and the method of consecutive approximations. The tools and the possibilities of their use for solving specific problems related to mine design of an open pit have been broadly discussed. The chapter in which the authors relate to the specific problems of Gubin deposit development constitutes the main part of the paper. It depicts the assumptions on which the annual production and the total extraction of lignite during life-of-mine period have been based. The process of mine designing has been discussed, taking into consideration environmental, geotechnical, spatial, social and mining conditions. The criteria for choosing the location of the opening cut and the external dump have been mentioned. In the summary, the present state of mine design progress has been presented, together with some problems that will need to be solved during the next stages of the deposit?s development.

Aug 1, 2013

By Song (Bruce) Li

This paper presents a three-tier web-based GIS application that uses Java Server Page (JSP), Java Servlet and relational database to manage performance monitoring and reporting in the oil sands mining industry. It is a real-time Earth Structure Performance Monitoring System (ESPMS), which allows the user to record and store instrumentation monitoring readings as well as technical observations and instrumentation installation documentations into a complete relational database. The ESPMS introduces the advantages of precise field record keeping, due diligence, security, consistency and completeness while at the same time simplifying the task of reporting. ESPMS combines web-based Geographic Information System (GIS) mapping and data management to facilitate the collection, organization, retrieval, reporting and communication of the vast amounts of monitoring data related to mining and tailings engineering associated with site investigation and construction information. ESPMS utilizes web-based GIS technology to incorporate referenced aerial photographs, Energy, Mines and Resources (EMR) maps or AutoCAD drawings as a base map for the monitoring site. It dynamically creates and administers all instrumentation monitoring locations and reading frequencies. It enables the user to collect and manage the data associated with an unlimited number of reading events. It provides the designer with real-time monitoring data (what, why, when, how), thereby allowing them to make early adjustments to design protocols or procedures, or to take remedial action(s) regarding the subject earth structure under construction or after completion. It provides more cost-effective reporting and data sharing through the Internet. It improves security on data entry and eliminates unauthorized data changes. It allows better integration of graphics and illustrations with numerical data files. It is capable of managing the continually expanding workload resulting from new and larger structures in the oil sands or other mineral mining industries. The ESPMS facilitates easier data gathering for preparation of as-built and after-built or regulatory reports, etc. It enables quick location of the interested instruments and summarization of the performance of the instruments. It enables one to visualize performance data and QA/QC construction test results spatially on any given cross-section. Data can be exported to MS Excel for detailed analysis and to evaluate the stability of the earth structures.

May 1, 2004

By J. Elbrond

"This paper describes qualitatively the progressive are losses and rock dilution in a mining project from the geological modelling to and including the concentration. A simple theoretical model of the net present value of a deposit, assuming a log-normal grade frequency function, is used to find the pair of decisions on rate of production and cut-off grade, which maximizes the net present value. Various degrees of ore losses and rock dilution are then applied on this optimum, and the resulting decreasing net present values are calculated, as well as the increasing cost of metal and the increasing metal losses as a function of are losses and rock dilution.IntroductionDue to a variety of uncertainties, to inevitable lack of precision in deposit estimation and in mine planning, and due to production constraints, ore losses and rock dilution occur throughout the many phases of a mining project. The phases range from the geological definition of the orebody, over the cut-off grade optimization and decision, over the design of the mining method and the actual mining to the ore dressing, where the final loss of valuable ore into the tailings occurs. Quantification of ore losses and rock dilution are mostly difficult if not simply impossible to obtain. At the end-phase an estimate of the total dilution may be obtained with some degree of precision by observations of the amounts and analysis of concentrate and tailings. It is much more difficult to estimate the total ore losses because the uncertainty of the geological modelling in the initial phase and in the mining phase is very difficult to quantify, despite the use of geostatistics for grade estimation.Many attempts at quantification of ore losses and rock dilution have been made and some examples are shown in Table I. Although they differ widely from each other, it is obvious from these figures that ore losses and rock dilution are significant and have considerable effects on the economical results of a mining operation. Life length, net present value, cost of producing metal and the loss of metal are all affected."

Jan 1, 1994

"Chairman - To start this discussion off, I wonder if Mr. Dave Livingstone or Mr. Jeffrey of Granisle Copper would give a description of the crushing plant at Granisle since it is somewhat different from the usual plant, W. N. Jeffrey - Granisle Copper The main difference between the Granisle crushing plant and many others is the Livingstone method of putting the screens under the crushers instead of above them. This method allows a more compact arrangement and eliminates the heavy superstructure required to take care of the screen vibration when the screens are mounted above the crushers. There is a small surge bin above the tertiary crusher which takes care of the variation in circulating loads.All three crushers are gyratory crushers. There is no grizzly ahead of the primary crusher but the fines are removed from the primary crusher discharge at the first screen and sent directly to the fine ore bins. E, Thornton - Sullivan GroupDo you have any coarse ore storage outside?w. N. Jeffrey - Yes, there is about 6000 tons live storage, We have very little trouble with freezing even though it is open storage.E. Thornton - Sullivan GroupDoes rain cause any problem?w. N. Jeffrey - It sometime creates a problem because the storage area is on a slope. It is mainly a problem of keeping the drainage ditches open.Question – What is the size of the primary gyratory crusher and how does it work on run of mine size materials, in other words, do you have a problem with oversize material and how do you handle the oversize rock?w. N. Jeffrey - The primary gyratory is a 42 x 65 with a nominal open side setting of 61'. We do at times get a run of oversize material. For anything too large for the crusher to handle, we use what is known as the Esco hook, which is a type of grapple used to remove oversize rocks.The disadvantage with this equipment is that the crane is too slow. I am sure there must be better arrangements than the one we have,"

Jan 1, 1971

By A. Miura, R. Winarko, D. B. Dreisinger

Chalcopyrite leaching in acidic ferric sulfate media at ambient temperature exhibits slow kinetics and poor leaching efficiency, which is typically attributed to surface passivation. Recently, the presence of small amounts of iodide in the leaching media has been reported to significantly accelerate chalcopyrite dissolution. To further explore the behavior of chalcopyrite leaching in the presence of iodide, it is essential to determine the factors that control the leaching kinetics. In this study, we investigated the effect of initial solution potential, iodide concentration, particle size, and temperature on the kinetics of copper extraction from chalcopyrite. The key in the experimental setup was to minimize iodide loss by iodine evaporation and precipitation as cuprous iodide. The former was achieved by performing the leaching tests in fully sealed leaching systems. The latter was achieved by maintaining the solution potential higher than 420 mV (versus Ag/AgCl electrode), where formation of CuI precipitate is possible. The experimental results show that higher solution potential and higher iodide concentration had a pronounced positive effect on the rate of chalcopyrite dissolution. There was no effect associated with particle size, which might be due to the narrow size range tested. The positive effect of increasing temperature only manifested itself when the solution potential was sufficiently high. INTRODUCTION The main challenge of chalcopyrite leaching using ferric ions as the oxidant at ambient temperature is the slow kinetics and poor extraction. The reason for this leaching behavior has been attributed to passivation resulting from formation of insoluble passivating layers on chalcopyrite surfaces, which could be metal-deficient sulfides, elemental sulfur, polysulfides, and iron precipitates (Klauber, 2008). To overcome passivation, technologies using higher temperature and pressure have been developed to leach chalcopyrite concentrates, such as the CESL process (Mayhew, McCoy, Mean, & Woeller, 2013), FLSmidth Rapid Oxidative Leach (ROL) process (Mulligan et al., 2017), Galvanox (Dixon, Mayne, & Baxter, 2008), Albion process (Hourn, Turner, & Hourn, 2012), HydroCopper (Hyvärinen & Hämäläinen, 2005), TPOX (Dreisinger, 2006), and AAC/UBC process (Dreisinger, Dempsey, Steyl, Sole, & Gnoinski, 2003). Even though they are effective in accelerating copper leaching kinetics, those technologies typically have higher capital and operating costs associated with production of concentrates and application of autoclaves. This makes them unsuitable for treating low grade primary copper sulfide ores.

Jan 1, 2019

By I. Silin, M. Stapelmann, H. Wotruba, K. M. Hahn, C. Gerold, J. Katzmarzyk

New deposits are often characterized as low-grade and complex ore bodies. To date, processing of fine grained ore deposits implies complex processing routes with high energy consumption. For that reason, the investigation of dry grinding, e.g. with vertical roller mills (VRM), is an approach to save energy reserves. It is already proven that dry grinding can influence subsequent processing steps. In terms of flotation, the surface characteristics of the minerals are crucial for a successful flotation process, e.g. the sufficient adhesion of mineral particles and air bubbles. Wet grinding results in different surface properties of the product entering the conditioning stage than dry grinding, which has direct influences on the subsequent flotation process. To investigate the consequences for flotation, processing of a copper sulfide ore containing chalcopyrite and bornite was investigated, including grinding and rougher flotation. The effects of wet tumbling mill and dry VRM grinding on the degree of liberation of valuable minerals are characterized using mineral liberation analysis by scanning electron microscopy. The obtained flotation results, as well as the recorded flotation parameters such as dissolved oxygen, oxidation reduction potential and conductivity are presented and discussed. INTRODUCTION Copper production is mainly connected to porphyry copper deposits as main source for copper sulfides such as chalcopyrite and bornite. State-of-the-art processing of these deposits is already established worldwide for decades, combining crushing, fine grinding and subsequent flotation with rougher, cleaner and scavenger stages (del Villar, Desbiens, & Maldonado, Bouchard, 2010). While the demand for copper is increasing, copper grades of the deposits are decreasing and the remaining ore bodies have more complex mineralogy and finer particle size distribution. This implies processing routes with high water and – due to the liberation of fine grained valuable minerals – high energy consumption by fine grinding. One of the main energy consumers in processing plants is the grinding section, which accounts for more than 50% of the energy consumption in processing plants and even 4% of the worldwide energy consumption (Jeswiet, & Szekeres, 2016). Therefore, dry fine grinding alternatives are applied to enhance the overall energy consumption of mineral processing plants: high pressure grinding rolls (HPGR) and vertical roller mills (VRM).

Jan 1, 2019

By Carlos Nieble

Ferteco Mineracao, belonging to the CVRD group, is one of the several iron ore mines in the south east of Brazil. Some of these mines have pit walls of over 350-m height and operations under difficult geo-technical conditions are steadily increasing. In order to mine the 15 million tonnes of high grade hematitic ore remaining in the Fabrica pit, Ferteco had to deal with the following restrictions: Warrant power to its concentrator and pellet plant, relocate the load-out rail road loop, stabilise a poor phillyte slope, relocate a creek on a bench, lower the water table and mine at maximum angles at a slope with phillyte and loose itabirites directly under the rail road loop. Considering these aspects, 3 stages of operations were designed, considering a safe operation, but with the lowest possible security factors at the last mining stage, before the pit would be filled in again with tailings. Stage one above the rail road loop and below the high tension power mast, where the poor phillyte slope was secured by a flexible retaining wall, using old rails and anchors up to 27 meters (80ft) with a working pressure of up till 30 tonnes and stage two with different slope angles, and deep wells, high wall blasting perpendicular to the highwall with relief holes and rerouting the creek in steel half culvert and stage three, the filling in with tailings and recuperation of the deep well pumps. Following studies and analysis were made: Geotechnical and geological studies of all rock formations including rock strength test, a hydrological study to define the lowering of the ground water level, slope stability analysis for every bench and all possible combination of benches, definition of blasting parameters close to the pit slope, a monitoring plan, including topographical movements, vibrations due to blasting, piezometers and pressure cells as well as a plan how to fill up the pit again with tailings in combination with the rising of the ground water level. A close co-operation between the geo-technical consultant, pit operations, the blasting contractor and the geological department resulted in an operation without any incidents, lowering the pit with an additional 3 benches and mining an additional 5 million tonnes of high grade ore.

May 1, 2003

By Benoît Plante

La mine Tio (QIT Fer et Titane Inc.) exploite à ciel ouvert un gisement d?ilménite massive depuis 1950, générant jusqu?à maintenant plus de 70 Mt de roches stériles potentiellement génératrices de drainage neutre contaminé (DNC), dont une partie est submergée dans le lac Petitpas, au nord de la mine. Si généralement l?ennoiement des rejets miniers a un effet bénéfique pour limiter la contamination des eaux, cela reste à prouver pour ce qui est des stériles générateurs de DNC. Des essais en réservoirs ont été réalisés afin d?évaluer la réactivité de trois différents types de stériles submergés, qui ont été choisis en fonction de leur teneur en ilménite. Afin d?évaluer l?impact de la matière organique sur la réactivité des stériles submergés, chaque type de stérile est étudié en immersion simple ainsi que combinée à des sédiments prélevés au lac Allard, à proximité de la mine, d?où l?eau utilisée dans les essais a aussi été prélevée. L?évolution des teneurs en Ca, Mg, Si et Ni durant les 8 premiers mois d?étude démontrent que les stériles immergés sont réactifs, mais l?impact de la matière organique demeure pour l?instant difficile à préciser. Tank tests on Tio mine waste rock The Tio mine (QIT Iron and Titanium Inc.) exploits a massive ilmenite ore as an open pit since 1950, producing over 70Mt of potentially contaminated neutral drainage (CND)-generating waste rocks. While the underwater disposal of mine wastes generally show a beneficial effect on water contamination, this method still needs to be approved for CND-generating waste rocks. Tank tests were realized in order to evaluate the reactivity in underwater conditions of 3 different waste rock types, chosen with regards to their ilmenite content. In order to evaluate the impact of organic matter on the reactivity of underwater waste rock, each waste rock type was studied in simple immersion and in combination with sediments sampled at Allard lake, located near the mine, where the water for the tests was also sampled. The evolution of Ca, Mg, Si and Ni during the 8 first months of testing show that the waste rocks are reactive underwater, but the impact of the organic matter is till hard to define.

Nov 1, 2008

By J. Dupre, A. Gontcharov, H. Pepin

Produced by the conventional TIG welding CMSX-4 welds had been studied. It was demostrated that microstructure of welds depended on filler materials, welding parameters and joint design. Epitaxial solidification of CMSX-4 clad deposits produced a directionally solidified structure while welding with the similar filler material. Defects free clad welds were obtained by TIG welding without preheating using similar and dissimilar filler materials.

Jan 1, 2006