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By B. R. Maas

Processing material at a consistent rate through the continuous countercurrent decantation (CCD) thickeners presented a major bottle neck to unlocking additional value from one of the post-flotation streams at the Kansanshi processing plant. The manual operation of the thickeners making up the CCD train resulted in frequent imbalances and process upsets across units. The automation of CCD trains is notoriously tricky as the available manipulated variables are outnumbered by the parameters requiring regulation. The implementation of an advanced, auction-based, control philosophy to manage the movement of both slurry and solution inventory through the train addressed this parameter imbalance and lead to the stabilisation of the circuit. Since the commissioning of this control strategy: an additional 7 200 dmt/day (+36% throughput) of material can reliably be processed through the train, the number of monthly downtime events has reduced, and a saving of $354 000/month in reagent cost has been achieved. INTRODUCTION CCD operation Agitated leaching is a common method used in hydrometallurgy to extract a metal from its host ore. The host ore is initially finely ground to liberate the surfaces of the valued minerals and improve the kinetics of the leaching process. The ore is then contacted in an agitated tank with a lixiviant to dissolve the desired metal into the aqueous phase. The resultant slurry from the leaching process comprises of a solution pregnant in the metal salt and a barren ore. The first purification step taken after leaching is to wash the pregnant liquor solution (PLS) from the depleted ore. On the industrial scale, a set of between two to ten thickeners arranged in a countercurrent configuration is often used to carry out this separation (Perry & Chilton, 1973). The slurry is introduced at the head of the train and a wash solution at the tail. The two streams are pumped through each thickener countercurrent to one other. At each stage the residual PLS in the slurry is mixed and diluted down with the countercurrent overflow solution from the downstream thickener. The diluted slurry is fed into the thickener where the solids separate out through gravity sedimentation. The settled solids are channelled towards the underflow of the thickener by a rake mechanism and pumped downstream to the next thickener in the train. The outcome is the recovery of a PLS from the overflow of the lead thickener and a slurry of washed solids from the underflow of the tails thickener.

Jan 1, 2019

By Matthew Russell, Bobby Randhawa, Leopoldo Mariscal Samaniego, Jose A. Puente, D. Brosig, Paykan Safe

The La Caridad copper smelter, located in Nacozari, Mexico, have started a multi-year plan to reduce sulfur dioxide and particulate matter emissions into the environment. In 2014 the smelter completed a voluntary plant-wide SO2 process emissions inventory as part of the smelter’s proactive environmental stewardship policy. In turn, this triggered the development of several environmental projects to alleviate these emissions. A strategy consisting of improved capture efficiency of the sources with greatest impact has been followed to maximize cost-effectiveness. The reduction of secondary emissions from the Teniente Converter, as well as of primary and secondary emissions from the Peirce-Smith Converters have been key areas of priority. Additionally, a new system to capture and recover particulate matter from the Anode Refining Furnaces off-gas is also in the process of being implemented. This paper presents an overview of the work developed behind this effort and the smelter’s vision towards a greener future. INTRODUCTION The La Caridad copper smelter is owned and operated by Metalurgica de Cobre S.A. de C.V. (a subsidiary of Grupo Mexico). Currently, the smelter has a production capacity of 300,000 tonnes of anodic copper per year. The smelter is coupled with a 300,000 t/y electrolytic refinery. Other by-products (such as precious metals and sulfuric acid) are also generated as part of an integrated process. The original smelter started operating in 1986 with one Furukawa concentrate dryer, one Outokumpu Flash Smelting Furnace (FSF), two slag cleaning electric furnaces (SCFs), three Peirce-Smith Converters (PSCs), two Anode Refining Furnaces (ARFs) and one casting wheel. In 1997, a plant expansion project was concluded with the installation of one steam dryer, one Modified Teniente Converter (MTC), one additional ARF, one additional casting wheel, and the electrolytic refinery [1]. Over the years, the smelter management has implemented numerous environmental projects to alleviate SO2 and particulate matter (PM) emissions. The smelter process gas handling system captures the process gases from the FSF, MTC, and the three PSCs. The process gases are cleaned and conditioned and sent to the two acid plants. The first acid plant (with a capacity of 226,000 Nm3/h, wet basis) was commissioned in 1988 to process the FSF and PSC off-gases. A second acid plant (with a capacity of 195,000 Nm3/h, wet basis) was commissioned as part of the smelter expansion project [1]. The PSCs also have a secondary gas handling system with a dedicated baghouse and stack. The Furukawa dryer off-gas reports to a dedicated electrostatic precipitator (ESP) for dedusting before venting via a local stack. Process gas from the two SCFs, along with hygiene ventilation from various sources inside the smelter (including matte and slag tapping hoods from the FSF) report to a baghouse-based gas handling system and vent to atmosphere via a local stack. The original ARFs do not have an off-gas handling system, however, each ARF has a post- combustion chamber which naturally vents the process gases to the atmosphere via a dedicated stack.

Jan 1, 2019

By Feng Tianran, Lu Shijie, Shen Zhengchang, Han Zhibin, Tan Ming

"As a kind of non-metallic material, high-purify quartz sand has been widely used in many fields. Flotation is a popular way to process quartz sand. But it is in difficult position for flotation to get high-purify quartz product due to large amount of coarse particles contained in the feed. In this paper, a design principle of special flotation cell for high-purify quartz sand has been put forward base on ore property analysis and laboratory test, and the pilot scale special flotation cell has been developed to meet the requirement of the quartz sand flotation process. The special flotation cell for high-purify quartz sand is featured by middle positioned impeller, false bottom and draft tube, which could decrease the rising distance of mineralized bubble and strengthen the circulation capacity of the lower part of the flotation cell. Hydrodynamic study indicates that high air dispersion degree and strong lower slurry circulation could be produced by pilot scale special flotation cell, which is helpful to create a suitable hydrodynamic environment for quartz sand processing.INTRODUCTIONNatural quartz is a non-metallic mineral, which is widely used as raw material in multiple industry process (Shifu, 2006). High purity quartz, especially containing more than 99.9% silicon dioxide, is characterized by high temperature resistance, small thermal expansion coefficient, highly insulating and so on. The demand of high purity quartz has been rising along with social and economic development, now high purity quartz is widely used in the field of photovoltaic industry, electronic information, communications and aerospace industry.Quartz is a kind of main rocking-forming minerals in the nature. There are seven genetic types, such as natural crystal, quartz sandstone, quartzite, vein quartz, konilite, natural siliceous sand, granite quartz (Baolei, Shuilin, & Dianchao, 2012). The natural crystal is primarily used as raw material to refine high purity quartz, but it is not available for two reasons. On the one hand, the natural crystal is extremely expensive and poor in resources reserves, which can not satisfy the need of large-scale industrial production. On the other hand, the chemical component of natural crystal is instability, it may lead to low stable product performance in large industrial applications, and it is also difficult to achieve raw materials standardization. Therefore, there are a lot of studies to search for other technological methods to obtain high purity quartz from other genetic quartz types."

Jan 1, 2016

By A. Kozlowski

We present research on cable fault location methods for mines where coal dust and methane explosions are a frequent potential hazard. The methodology and devices for low- and medium voltage cables fault location are well known and applied in classical power distribution networks. They are based on solutions for underground mines with a high risk of explosion. The lack of a safe solution for the location of a fault has contributed to the initiation of a research project by a group of Polish companies. The consortium consists of research institutes (Institute of Innovative Technologies EMAG, Tele and Radio Research Institute, Central Mining Institute GIG-KD ?Barbara?), Silesian University of Technology and a private entrepreneur MARTECH PLUS. Fast and exact determination of fault location is one of the important factors for minimizing power outage time. The lack of technical solutions to obtain measurements in a secure way leads to the use of the methods connected with a high risk of explosion. For example, the application of high energy can lead to arc discharge where there is a fault in insulation, creating the risk of an explosion. Measurements of methane concentration along the analyzed section of the cable reduce the risk; however the shock wave generated by the electric arc may raise coal dust and make an explosive mixture, which is an object of ignition. Low energy methods, like those based on reflection waves, have low sensitivity for high- and medium-resistance faults. This technique has difficulties in joints, which create false echoes. The fundamental problem is to determine the topology of the network using a reflectogram and find anomalies of a different nature ? not only open and short circuit. Included in the scope of the project is an analysis of Polish legal regulations related to fault location and operational work for low- and medium-voltage networks. Two of the methods used on the surface were selected as feasible to implement under explosion hazard conditions. In the process of development of the fault location methodology, installation features characteristic of underground mines were used, namely access to the whole length of a cable. This is the attribute that provides an opportunity to do measurements that are usually not possible for typical surface power nets. To carry out laboratory tests, a dedicated measuring stand with various types of cables was built. The authors describe the principle of operation of the fault locator and the results of the laboratory tests carried out on the cables with built-in damage (prepared for tests). Results of the work will be the basis for the next stage of the project associated with operational analyses of the prototype.

Aug 1, 2013

By Alan B. Morrow, William A. Poe

"Minerals processing operations are constantly challenged with improving production rates, yields, energy efficiency and process stability. Advanced Process Control has proven to be an effective tool for improving operations by capturing the full process capabilities consistently. However, the velocity of applying the technology as well as the sustainment of these significant benefits has been hampered by the usability of the software and adaptability of the applications.Some of the new technology developments that improve the usability and adaptability of advanced process control include:• modern and intuitive user interfaces,• customizable models to support unique process behavior,• fast and flexible OPC connection tools,• automated model building,• integration with first principles dynamic models and• improved performance monitoring.This paper will briefly discuss the new technology developments that have improved the usability and adaptability of advanced process control and how this can affect minerals processing operations.In today’s environment, Minerals Processing plant managers continuously face new challenges and requirements including:• higher energy efficiency,• increased run lengths,• reduced product quality variability,• reduced emissions,• higher availability, and• lower operating costsOperating a unit to meet these challenges in real time is not easy for operators and conventional process control systems. However, innovatively using existing technologies in Minerals Processing plants can improve operations on a minute by minute basis and reap significant benefits.Advanced Process Control (APC) can extract the maximum performance benefits from existing equipment while respecting plant operating constraints. Expensive capital projects to upgrade equipment are not always necessary. Advanced Process Control can deliver significant economic benefits by maintaining the plant at its most profitable point as conditions change."

Jan 1, 2016

By F. Osmond, R. Kelly

"Malcolm Knowles, the famous adult educator once wrote: ""What a challenge we in human resources development face if we are to avoid the obsolescence of our workforce. I can foresee this challenge requiring that we reconceptualize (...) and redefine the role of HRD, away from that of managing the logistics of conducting training activities to that of managing a system of learning resources"".The context in which this was written surrounds the subject of adult education and the need for business and industry to change its approach to the treatment of adults as learners. In explaining what 'managing a system of learning resources' means, Knowles refers to line supervisors and managers as the most ubiquitous resources for day-in and day-out employee development. In many organizations training has been largely a staff rather than line responsibility. Even where trainers reported to line management, it was still seen as a staff function performed by a few specialists.This change referred to by Knowles places a much heavier responsibility on line supervisors and managers for the development of their personnel than has traditionally been the case. Managing that change - handing over the responsibility to the line organization - is the supreme challenge for the HRD function.Just why is this change necessary? Because too often, the learner does not effectively use on the job what he or she learned in the classroom. In fact, learners often discard new methods and revert to the older ways. The reason training does not stick is that the systems for training and work management are most often separate entities and while one may support the other, they need to be a single integrated system in order to ensure the transfer to the job effectiveness of training and some real value for the training dollars.Because much training can be wasted, steps have to be taken to ensure that it fulfills its intended purpose. The effective and continuing application of knowledge and skills that learners acquire can only be realized when line supervisors and managers are at the centre of the training system.This paper describes how such a systemis being developed at the Iron Ore Company of Canada (IOC) as a functional training program is introduced. The following text provides background information on the IOC - who they are and where they are located, a definition of functional training, an evaluation of the suitability of this approach for adults in industry, a description of how the program is being implemented at IOC and some observations that have been made as this change is experienced."

Jan 1, 1995

By Takao Oishi, Go Ohara, Toru Kitazaki, Masahiro Tanaka

"The High Pressure Acid Leach (HPAL) plant for nickel-cobalt mixed sulphide production at Coral Bay Nickel Corporation (CBNC) in the Philippines has been in operation since 2005 with a capacity of 10,000 t/year of nickel. Subsequent to the commissioning of the second plant in 2008, the production capacity increased to 22,000 t/year of nickel (Llerin et al., 2011). CBNC’s annual nickel production reached more than 24,000 tons of nickel in 2014. The concepts to achieve this capacity are the high availability and the reciprocally flexible operation of both plants. The availability depends on some factors, such as the control of impurities in the process and the equipment condition, to mention a few. In the HPAL process, it is not only the autoclave operating conditions that affect nickel extraction but also the impurities in the ore greatly influence it. The major element in the ore is iron and most of this iron is precipitated as Fe2O3, while the rest remains as Fe2+ or Fe3+ sulphate in solution. These ions in solution reduce the availability, so it is important to control the formation of iron ions by controlling the impurities in feed ore as well as the HPAL conditions. This paper presents the general HPAL process and the behaviour of impurities, focusing on iron behaviour in the HPAL process.INTRODUCTIONCoral Bay Nickel Corporation ProfileCoral Bay Nickel Corporation (CBNC) was founded on April 2002, as a joint venture of four companies; namely, Mitsui Corporation (18%), Sojitz Corporation (18%), Nickel Asia Corporation (10%) and Sumitomo Metal Mining Corporation (SMM) with a controlling stake of 54%. Aside from its capital share, Nickel Asia Corporation also supplies the ore and limestone for the High Pressure Acid Leach (HPAL) plant located in Rio Tuba, Philippines, as shown in Figure 1 (Tsuchida, 2008). The objective of the CBNC project is to produce 22,000 t/year (original design) of nickel and 1,500 t/year of cobalt, as sulphide, by the treatment of nickel laterite ore which is a low grade nickel oxide ore in the Philippines (Mines and Geosciences Bureau, 2014). The sulphide product has been exported as a raw material to the SMM Nickel Refinery in Japan for 30 years. The product of this plant, a mixed (Ni/Co) sulphide, is shipped to the SMM’s Nickel Refinery in Niihama, Japan. SMM Corporation provides the technology and management personnel to the plant. SMM Corporation also supplies the sulphuric acid, which is the main reagent for the HPAL process."

Jan 1, 2016

By Z. D. Hora

Several deposits of crystalline magnesite are known in southeast British Columbia. Deposits at Cross River and Marysville have been thoroughly explored, but as yet there has been no commercial production.

Jan 1, 1984

By Leopoldo Gutierrez, Sergio Castro, Lina M. Uribe, Janusz S. Laskowski

"The effect of sulfite and metabisulfite ions on the amyl xanthate flotation of copper-activated pyrite, in distilled water and seawater, has been studied by micro-flotation tests and redox potential measurements. The activation of pyrite was carried out varying copper ions concentration. It was found that in the presence of copper ions in solution, sulfite and metabisulfite are able to depress copper-activated pyrite at neutral or slightly acid pH. After the addition of sulfite or metabisulfite ions the resulting pH is more acidic with metabisulfite ions. A significant increase of the redox potential of the solution was observed as copper ions concentration increased. Then, the initial reducing conditions brought about by sulfite and metabisulfite ions, change abruptly to oxidizing conditions in the presence of copper ions. Under these oxidizing conditions, the metabisulfite showed to be an effective depressant for pyrite mainly in seawater. However, sulfite was a pyrite depressant in both distilled water and seawater. Based on these results, the pyrite depression by sulfoxy-depressants during the flotation of Cu/Mo ores in distilled water and seawater is discussed.INTRODUCTIONDepression of pyrite by sulfoxy-depressants (sulfite and metabisulfite) in the presence of copper ions is an important issue. This is particularly true when pyrite depression in the flotation of Cu/Mo sulfide ores is carried out in seawater. The use of lime in seawater must be avoided because the formation of magnesium hydroxy-complexes and precipitating magnesium hydroxide strongly depresses molybdenite flotation (Castro et al., 2014). In order to solve this problem, one large plant in Chile uses sodium metabisulfite (MBS) instead of lime to depress pyrite in seawater. In plant practice, this also improves molybdenite flotation since with the use of MBS the pyrite depression is carried out at pH 6-7, instead the pH 11-12 usually employed with lime.From a general point of view, pyrite exhibits some natural floatability over acidic pH ranges and may be partially floated without thiol collectors (a coating of sulfur/polysulfides has been claimed). Therefore, pyrite flotation with xanthates is usually high in the acidic pH region, but decreases greatly at alkaline pH (Leppinen, 1990; Fornasiero and Ralston, 1992; Zhang et al., 1997)."

Jan 1, 2016

By B. McFadzean, I. Achaye, T. Chidzanira, J. Wiese

"Froth stability has an overriding effect on metallurgical recovery and grade for many ore types. Froth recoveries of less than 10% are common for low stability froths such as those in sparsely mineralized ores or in cells at the back end of flotation circuits. It is well known that, aside from chemical factors such as frothers and electrolytes, the mineral particles themselves have the dominant influence on froth stability. The particle properties of size and hydrophobicity are expected to have the greatest effects on froth stability. The aim of this study was to understand the effects of particle size on froth stability and to quantify these relationships.The study compared a synthetic ore of pyrite and quartz, with carefully controlled size ranges and hydrophobicity, with conventionally milled platinum-bearing UG2 ore and the bulk reverse flotation of hematite ore. A novel bench-scale continuous column flotation cell and continuous mechanical hybrid cell with large relative froth volumes, as well as a three-phase froth stability column were used to characterize the froths.It was found that froth stability could be related to the concentrate particle surface area by a Langmuir-type relationship. This is indicative that the mechanism by which particles stabilize the thin films in froths is similar to the mechanism by which surfactants lower surface tension. Different froth stabilities were found for different ores and this was thought to be the result of different particle sizes and distributions, which resulted in different packing densities in the thin films and Plateau borders. Interestingly, there was a clear relationship between the feed particle size and the froth stability even though froth stability was dependent on the concentrate particle surface area. It seems that the feed particle size is a good predictor of the surface area of particles in the concentrate. Concentrate particle size did not show a clear relationship with froth stability since it was already shown that it is both the size and the amount of the particles in the froth that contribute to the stability.INTRODUCTIONThe stability of the froth in froth flotation is important to the recovery and grade of the valuable mineral. Too stable a froth and the grade is reduced; not stable enough and the recovery may be compromised. Platinum-bearing ores with fragile, poorly mineralised froths are known to have froth recoveries that are routinely less than 10% (Crosbie et al, 2009). There are a number of chemical, physical and operating parameters that can affect froth stability, including reagent dosages, air flow rate, pulp density and particle properties. Arguably, the largest contribution to froth stability is from the particles that stabilise (or destabilise) the thin films in froths. Particle size, hydrophobicity and shape all affect froth stability to some extent."

Jan 1, 2016

By Bernard Tougas, Franco Chiesa, David Levasseur, Bernard Duchesne

"In the present paper, the optimization of a tilt poured permanent mold will be carried out using filling and solidification modeling. This exercise resulted in 3 modifications of the initial mold design which was based on the foundry experience. The optimization provided a 20% improvement in the yield while increasing the output from 10 to 15 castings per hour.In a subsequent study, the value of the critical fraction solid of AlSi7Mg (A356) alloy was verified by radiographic analysis of the castings produced in the first 5 cycles of a run, which are necessary to bring the mold to a dynamic thermal equilibrium."

Jan 1, 2014

By R. G. McDonald

"High iron solution concentrations often result in the formation of basic ferric sulfate during the high temperature oxidation of sulfidic materials such as base metal and refractory gold concentrates. Consequently, the “Hot Cure Process” followed by limestone neutralization is required to facilitate subsequent processing and/or disposal of the leach residue. The present paper discusses strategies that can be employed to influence the composition and nature of the high-iron content leach residues produced during pressure leaching.INTRODUCTION Control of iron deportment in hydrometallurgical processing remains a key challenge for a range of reasons. Strategies generally relate to the control of conditions under which iron is optimally rejected (Demopoulos, 2009). Supersaturation provides the driving force for crystallization and this can be controlled by a number of factors that include (1) pH and/or Eh control, (2) the presence of matrix anions and impurities, (3) dilution, (4) complexation and dissociation and (5) manipulation of the rate at which the metal to be precipitated is released to solution. Each of these factors is impacted by temperature. Good control of crystallization is essential to the control of precipitate properties such as the degree of crystallinity, stability, particle size and quality which covers both the extent of impurity uptake and suitability of the precipitate for waste disposal. The quality and properties of iron precipitates in turn can impact downstream processing. It is the aim of this paper to discuss the generation of iron-containing residues produced during the pressure oxidation of sulfidic ores and concentrates and how the nature of these can be controlled to potentially facilitate further processing. A brief overview of several related topics will now be given and include, where relevant, discussion of work undertaken in our laboratories. In addition, brief discussions of selected process options that enable good control of iron rejection will be provided."

Jan 1, 2016

By Avi Nanda, Robert Brandt, Alex Piatkiewicz, David M. Jones

In 2010 the United States Environmental Protection Agency (UESPA) revised the National Ambient Air Quality Standards (NAAQS) by lowering the permissible value to a 1-hour average SO2 concentration of 75 ppb. In 2012, the USEPA designated the Miami area as non-attaining for the new SO2 NAAQS. A number of options were studied to assist in bringing the Miami smelter into attainment with the new standard. The chosen solution included: Installation of a large tertiary hooding system over the converter aisle; Hooding of the anode vessels and treatment of the anode gas in a baghouse; Installation of a large 2,325,212 Nm3/hr caustic scrubber to treat the captured converter fugitives and anode off-gases; Installation of a secondary ESP for the tapping fugitive off-gas treatment system; and the continuous scrubbing of all acid plant tail stack gases. The modifications have reduced smelter SO2 emissions by 85%, achieving a world class SO2 stack emissions rate of 0.45 kg of SO2/tonne Cu. INTRODUCTION The Freeport McMoRan Miami smelter has been previously described in literature (Steinborn, Pasca, Sallee, Binegar, Bhappu). For the sake of brevity a brief description is given here. The Miami smelter is located 129 km east of Phoenix, Arizona near the town of Miami, Arizona. The smelter was established in 1915 and has undergone a number of process upgrades and modernizations over the years. The present smelter consists of an Isasmelt furnace, an electric settling furnace, four Hoboken converters, two anode refining vessels & one casting wheel; and one double contact sulfuric acid plant. Concentrates are railed & trucked to Miami from four Freeport McMoRan mines located in Arizona & New Mexico. The wet concentrates, typically containing 10% H2O with a Cu/S ratio of 0.85, are smelted in an Isasmelt furnace to a matte grade of 58%. The entire output from the Isasmelt is laundered into an Elkem electric furnace (originally installed in 1972 as the primary smelting furnace) for separation into matte and slag. Slag containing 0.85% Cu is skimmed from the electric furnace and discarded. Molten matte, containing 60% Cu, is tapped from the electric furnace and processed in one of four Hoboken converters to produce blister copper. Two anode refining vessels further refine the blister copper before the final product is cast on a single, 30 mold, casting wheel. All of the anodes produced at Miami are shipped to the Freeport McMoRan electrolytic refinery in El Paso Texas for refining into cathode copper. Smelter off-gases from the Isasmelt, electric furnace and Hoboken converters are combined and treated in a double contact acid plant to produce sulfuric acid. Acid produced at Miami is internally consumed at Freeport McMoRan mines in the region.

Jan 1, 2019

In the last 30 years numerous mining plants have been developed in North America but only a handful of these have required freezing the area where the shaft was sunk. The most recent mine shaft to be sunk using this technique is located on an island off the coast of Louisiana. The project required the input and skills of numerous people in several companies, including the owner, engineers and contractors. The Cote Blanche installation is one of several mines in the area which extract salt from a dome which extends under the Gulf of Mexico. The geology of this area is, however, markedly different from others in the region because there is approximately 460 ft of water logged soil above the salt dome. Because it is not possible to sink a shaft in ground that is this soft, the overburden where the shaft is located must be frozen before excavation can take place. The Cote Blanche mine is a producing facility which was built in 1958 by the Carey Salt Company. The original facility included a 14 ft diameter production shaft and a 7 ft-8 in diameter service/ventilation/escape shaft. This system was installed by The Cementation Company. In 1996 the owner decided to improve both the mine ventilation and production hoisting system by installing a third shaft. This shaft would be designed to extract 650 tons per hour from a depth of 1600 ft and provide 500,000 cfm of ventilation to the mine at a maxium design pressure of 4 in WG. These parameters resulted in a requirement for a 16 ft finished inside diameter concrete lined shaft and a headframe designed as a pressure bulkhead. Because the mine would operate from one level only, the most efficient hoisting system for the application was determined to be a ground mounted koepe running on rope guides. With these criteria fixed Cote Blanche embarked on a program which would see construction start in June 1998 with the first salt produced in February 2001 at a cost of $US18M. The major players in the design and construction were G.L. Tiley & Associates Ltd., BLM Bharti Engineering, Freezewall Inc., Phillips Mining, Geotechnical and Grouting Inc., Don Hart Consulting and Dynatec International.

May 1, 2002

By Zhongxue Li

This paper reviews an analytical model that defines the relationship of weight-weighted or volume-weighted average haulage distance with bench height and other related parameters, presents the optimum bench heights under various conditions in terms of minimizing the weighted average haulage distance of ore/waste during the full lifetime of an open pit mine, and discusses the truck haulage economics of mining open pits by high bench, specifically by 15 m, 18 m and 20 m benches through marginal analyses. The observations from these analyses disapprove the previous conflicting assertions that the over-all weighted average haulage distance either increases or decreases directly with bench height. It is concluded that: (1) an increase of bench height to 15 m from 12 m is likely to be viable under the general conditions of large-scale deep open pit mines; (2) the reduction of haulage cost is much limited with the case of increasing bench height to 18 m from 15 m; and (3) the feasibility of mining open pits by 20 m benches is generally questionable. These conclusions may serve as a useful guide for those who advocate open pit mining by high bench.

Jan 1, 1995

By Ngonidzashe Chimwani, David Glasser, Diane Hildebrandt

"Milling circuits play an important role in achieving the desired product; tonnage, size distribution and properties, since they are directly linked to the comminution energy consumption for the whole process. A grinding model with a mathematical simulation was developed for an open milling circuit with a scalped feed from a hydro-cyclone. The circuit was optimized by determining the optimal conditions in terms of energy consumption which would produce a floatable product having a desired size range (i.e. -75 +9 µm) for a platinum-bearing ore. The simulation results showed an operating region comprising the best combination of the ball filling and mill speed to achieve the highest production of a floatable size class, with the minimum energy consumption.INTRODUCTIONBall milling is usually conducted in closed circuits in order to achieve a desired product size without material overgrinding. The technological importance of optimising milling circuits cannot be overemphasised (Toneva and Peukert, 2007). Previous work has brought significant insights into the benefits of modelling milling circuits (Austin et al., 1984; King, 2012; Toneva and Peukert, 2007). However, the ability to model these circuits is dependent on the performance of the classifiers and the fidelity of the grinding mill model used compared to the actual circuit (Toneva and Peukert, 2007). Austin et al. (1984) identified five types of milling circuits as shown in Figure 1 namely; the grinding circuits employing mill in an open circuit, normal closed circuit, reverse closed circuit, open circuit with scalped feed, and, the combined closed circuit. Depending on the presence of the first classifier, the second, or both as well as their corresponding loops (streams), the combined circuit (e) can reduce to; an open (d), a normal closed (b), an open circuit with scalped feed (c) or a reverse closed circuit (a). In the case of an open circuit, the mill is operated without being directly influenced by a classifier. The normal closed circuit does not have the first classifier and therefore the make-up feed is the actual feed stream. Conversely, the reverse closed circuit keeps only the first classifier. The open circuited with scalped feed has only the first classifier but does not have the second one. The overflow stream from the classifier is mixed with the mill product to form the final product for the downstream process."

Jan 1, 2016

By David Rahal, Saul Ortiz, Maria Escobedo

The general trend in the minerals industry is towards the processing of fine grained, disseminated mineral deposits. This has led to an increase in the number of fine/ultra-fine grinding installations using stirred media mills. The challenge is that both the equipment type and the mineralogy of the deposit have an effect on specific energy consumption. Fresnillo PLC recognized this issue and addressed it by providing samples for testing in a laboratory regrind system. The objective of the test program was to determine the specific energy required to achieve a product size of 10 µm (P80). A specific energy plot was generated for both a lead and zinc flotation concentrate. It indicated that 34 kWh/t and 51 kWh/t were required for the two samples (Pb and Zn respectively). This difference highlights the effect of mineralogy on the energy required for fine grinding. The regrind feed and products were submitted to a range of basic mineralogy tests to quantify the mineral differences that could have changed the specific energy requirements. These tests provided general mineralogy, documented optical textures, and provided particle size microscopy. The lead concentrate was found to contain significant amounts of silicates in addition to the dominant sulphide, pyrite. The zinc concentrate was determined to be mostly sphalerite with minor amounts of quartz, chalcopyrite, pyrite and galena. The material testing results illustrate the importance of basic mineralogy studies in conjunction with comminution testing to determine specific energy consumption.

Jan 1, 2016

By E. Ghali, W. Zhang, M. Robichaud, G. Houlachi

"The performances of six mesh and one plate dimensionally stable anodes have been investigated in the zinc electrolyte at 38oC and compared to conventional silver-lead alloy anodes. The electrochemical behaviour of these anodes was studied by cyclic voltammetry, galvanostatic, potentiodynamic, open circuit potential and electrochemical noise measurements. It was observed that at the end of 24 h of galvanostatic polarization at a current density of 50 mA/cm2 in the zinc electrolyte at 38oC, the mesh V had the lowest potential followed by mesh III, mesh VI, plate-V, mesh IV, mesh I, and mesh II electrode. When the potentials for mesh I, mesh III, mesh IV, VI and plate-V, mesh V were compared with that of the Pb-0.7%Ag anode (1920 mV/SHE), they were found to be 6%, 16%, 7% 10%, 17% lower, respectively.After 16 h decay following the 24 h polarization, the mesh III had the lowest corrosion rate (51 mS/cm2) among the seven specimens, followed by the plate-V, mesh V, VI, IV, I and II (221 mS/cm2). For the fresh surface as received, after two hour open circuit potential measurements, mesh I had the most positive (noble) potential (1382 mV/SHE) among the seven electrodes followed by the plate-V electrode, mesh V, mesh IV, mesh VI, mesh III and mesh II (751 mV/SHE). After consecutive series of electrochemical testing, mesh V, mesh III, mesh VI had lower corrosion rates (~ 343± 6 µA/cm2) than that of mesh IV, mesh I, plate-V, and mesh II (~ 644± 143 µA/cm2).INTRODUCTIONThe continued use of lead based anodes in the electrowinning process has been mainly due to their relatively low cost as compared to other materials (Msindo, Sibanda, & Potgieter, 2010). They have good electrical conductivity, stability during electrolysis, very good corrosion and wear resistance (Ivanov, Stefanov, Noncheva, Petrova, Dobrev, Mirkova, Vermeersch, & Damaerel, 2010). However, lead-based anodes give their high energy consumption and low corrosion resistance for electrowinning. To overcome these problems, Dimensionally Stable Anodes (DSAs) has been chosen (Msindo et al., 2010). Comparing to the traditional lead anodes, the DSA anodes have many advantages: (1) no contamination; (2) no loss of anode material, pure electrolyte; (3) lower gas bubble effect through special anode designs; (4) use of higher current densities (Msindo et al., 2010). The success of DSA anodes has been attributed to their chemical stability that allows them to maintain nearly constant dimensions during the life of the anode even when operating in environment with low pH values. Due to their stability, the consumption rates of DSA are as low as 1 mg/amp.y (Alfantazi & Moskalyk, 2003)."

Jan 1, 2016

By Mike McKevitt

We are trying to promote a concept and a change in management?s approach to the iterative process of grade control. This is important as grade control must routinely try and achieve the long term objectives set by the mine plan ??BackgroundBackground ??Factors contributing to variation from planFactors plan ??Basis of OptimisationBasis Optimisation ??Ore Block Optimisation processOre process ??Operating mine examplesOperating examples ??Chain of Mining. ??Intention is to highlight the need for repeatable and consistent materialand material-type classification to remain consistent with the Life of Mine Plan ??We can drill another hole or reWe re-optimise a plan, but once excavated, the final decision has been made ??The grade control decision is the last decision made in the resource exploitation sequence.

May 1, 2009

By Yuhua Wang, Wei Sun, Dongfang Lu, Yuehua Hu, Tao Jiang

"With the continuous depletion of high-grade tungsten ores, the low-grade refractory wolframite has become an important source for the tungsten industry. However, the large-scale utilization of low-grade refractory wolframite is disappointedly scarce, it is due to leanness of valuable minerals, complex mineral characterization and use of toxic collectors.The aim of this study was to develop an environment-friendly processing flowsheet, which was used for beneficiation of low-grade tungsten ore containing 0.45% WO3 from New Brunswick in Canada, and the content of arsenic is 0.68% in this materials. The objective of beneficiation studied was to obtain a concentration containing 60% WO3 with less than 0.5% of arsenic under the condition without any flotation reagents. In this environment-friendly processing flowsheet, magnetic separation technology, classification and gravity separation technology were combined and employed. Firstly, magnetic separation was employed for the first pre-concentration process, over 95% of the feed was rejected as the final tailings, meanwhile, over 85% of arsenic minerals were removed with tailings. Secondly, gravity separation tests were employed on the classified magnetic pre-concentrate in the second pre-concentration process, a second pre-concentrate with WO3 grade of 54.32% and As grade of 5.41% was obtained. Finally, two stage magnetic cleaning were conducted on the second pre-concentrate for upgrading the second pre-concentrate, the mostly arsenic and gangues were removed from the second pre-concentrate. Finally, this novel processing achieved an effective recovery of fine wolframite from low grade tungsten ore with high arsenic. A concentrate with WO3 grade of 64.81% were obtained in new opened processing flowsheet, and As grade was less than of 0.5% in this concentrate. Furthermore, remarkable feature of this processing flowsheet is environment-friendly due to avoid using the toxic flotation reagents during the whole mineral processing.INTRODUCTIONTungsten is an essential metal and is the most widely used in industrial production such as national defense, electro-mechanical products, textile and leather sectors (Srivastava et al., 2000). Wolframite and scheelite are the main mineral sources of tungsten metal. With the continuous depletion of high-grade tungsten ores and the increasing demand for tungsten alloy, the low-grade refractory wolframite has become an important source for the tungsten industry. However, the large-scale utilization of low-grade refractory wolframite is disappointedly scarce, it is due to leanness of valuable minerals, complex mineral characterization and heavy use of toxic collectors."

Jan 1, 2016