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By X. Q. Cao, D. S. Zhang, L. Q. Ma

To solve the support stability control problem for a coal face with large angles along both strike and dip (CLSD), the ’support-surrounding rock’ mechanical model has been developed, which takes into account the impact of the dip angle of the seam on the stability of the support in the strike direction. The mechanical relationships of the critical topple angle and critical slip angle of the support along the strike of the coal face with large dip angle and the support height, support resistance, friction coefficient, and other factors have been derived through the mechanical analysis of support stability in the strike direction of CLSDs in the free state, the operating state, and the special state. The research findings were applied to a fully mechanized CLSD in Xinji Coal Mine. The maximum underhand angle and overhand angle in strike are 42° and 25° respectively, and the maximum dip is 39°. It is calculated that during underhand mining and overhand mining, the critical support resistances for avoiding support toppling are 3723 kN and 1714 kN respectively, and the critical support resistances for avoiding slipping of the support are 7405 kN and 6606 kN respectively. Thus, the selection of type ZZ7600/18/38 hydraulic roof support for the coal face is justified. Measures to prevent sliding of the support and the installation of a limiting stop maintain the support runs in good condition and ensure safe and efficient mining of CLSD.

Jan 1, 2015

By G. Couture

"The Glencore Raglan mine Concentrator, located in Quebec’s Nunavik region, has been in operation since 1997. The nickel metallurgical plant was originally designed to process 100 tonnes per hour of ore to produce 130 000 tonnes of nickel concentrate per year at 16% Ni. The increased grinding capacity and the discovery of rich ore bodies have increased the annual production to more than 190 000 tonnes of concentrate at 19% Ni. In 2014, the Metallurgical group, with support from BBA, implemented an expert control system on the flotation circuit. This project was initiated to cope with the high variability of the concentrator’s feed. Increasing the plant recovery and final concentrate grade were also important project objectives. The implementation of the expert system ensures better control of concentrate inventory while maintaining maximum plant throughput. This paper describes the successful implementation of the new control system, the control technology selection and the flotation circuit optimization. It will also discuss the performances obtained with this new approach and the improved process stability.IntroductionThe Raglan Mine is located in Northern Quebec along the 62th parallel. It has been in operation since 1997 and was originally designed to produce 130 000 tonnes of concentrate annually at 16% Ni. The increasing of the grinding capacity and the discovery of rich ore bodies have increased the annual production to more than 190 000 tonnes of concentrate between 18 and 20% Ni."

Jan 1, 2015

By G. Song

"SynopsisAccording to the overlying rock strength and thickness of 10 shallow coal seams in China, the strata are classified into four types: weak and laminated strata (hereinafter referred to as WL), weak and massive strata (WM), strong and laminated strata (SL), and strong and massive strata (SM). The four types of strata, together with longwall panel height (M), panel length (L), and seam depth (H) are factors that influence the loads on hydraulic power supports. Principal component analysis was used to study the seven influencing factors. It turns out that seam depth (H) is multicollinear with WL, WM, SL, and SM. Thus the enumerated seven factors are divided into two models. In Model 1, six factors; panel height, panel length, and layer thickness of the four type strata are involved. In Model 2, three factors; panel height, panel length, and seam depth are included. Using multivariate regression analysis, two equations to calculate loads on supports are obtained for each model. The equations were applied to shallow coal seams. The equation of Model 1 is more accurate than that of Model 2, but Model 2 could be further used in deep coal seams. The equations were coded through VB programming to simplify the calculation.IntroductionTo ensure coal production in the long term, the focus of coal mining has been gradually shifted to the western areas of China in recent years due to rich coal resources in this region. Most of the coal seams in this area are shallow but the support loads are abnormally high. In these circumstances, the chock loads estimating methods used in the past do not work. If the rated working resistance of a support is larger than needed, the shield is unable to fully perform, while a lower working resistance would lead to accidents like shield crushing or even casualties. This problem is frequently found in shallow coal seams such as Shenfu-Dongsheng, Yulin, Tuha, Lingwu etc. (Zhang, 2011). Thus, to ensure safety and efficiency during mining, the estimation of support loads and selection of supports are of great significance."

Jan 1, 2015

By P. V. Chandramohan

"In ports, piled berths are designed and constructed for berthing of ships with a finite draught. Piles in a bridge foundation also have a large free stand above scoured level. The piles of these structures have to withstand large horizontal forces imparted by the impact of the approaching vessels or due to longitudinal and transverse action in a bridge. It is also imperative that these piles have to project out of the dredged / scoured bed. These piles are normally designed as slender compression members subjected to bending also. They are subjected to vertical load tests for assessing the soil resistance. Lateral load tests, as a rule, are also stipulated. IS and IRC codes stipulate a deflection criteria for lateral load test to arrive at the capacity. These tests are usually conducted on a working pile by jacking out two piles. At the time of the test, pile will be free headed. But in service, piles are fixed head. There is a large difference in deflections between a fixed head and a free headed pile. Besides, a free headed pile will be subjected to a much higher bending moment than the fixed head one. This will put the working test pile under risk of breaking during the test. The problem is analyzed with case studies. INTRODUCTION Vertical capacity of piles is assessed by static formulae. This formula is based on the soil properties obtained from random soil exploration and laboratory and in-situ testing. Since soil is not a homogeneous medium, properties can vary. The design parameters taken for computations may be different from the ones at the interface of the pile with the soil. The best way is to find out is by a load test. The code IRC:78-2014 has rightly stated that initial load test is part of the design process. This is true with the vertical load test. But there are certain problems with the purpose and inference of the lateral load test. In both the tests the criteria for acceptance is the quantum of movement of the point of application of the load. In the former, the movement is vertical and depends more on the compression of the soil than the compressive strain in concrete. But in the latter, this movement is partly on account of the flexural rigidity of the pile and partly due to the horizontal compressibility of the soil. The role of rigidity of the pile is more when compressibility of the soil is higher. Lateral load test is done by jacking one pile against another because it is difficult to get an independent horizontal reaction. Please see Fig.1 Photo."

Jan 1, 2015

By Madan Kumar Annam, Hari Krishna Y

"Basement construction is one of the challenging geotechnical concerns in the urban environment. Several types of retention systems are practiced to support the earth which is suitable for restricted access, rapid mobilization, fast excavation and cost-effective designs. A thorough understanding of geotechnical engineering as well as the structural mechanics is important for the earth retention systems. This paper discusses a project with various types of foundation solutions including monitoring of deflections during excavation stage of a retention system.A multi storeyed residential building (3B+Stilt+18F) in Chennai having structure foot print area of 1500 m2 has to be constructed on large diameter pile foundations and also to support the proposed basement excavation. Ground water table was encountered at about 6m below the existing ground level (EGL). In order to support heavy structural loads, large diameter bored cast-in-situ piles were chosen and installed. Further, complete retention system was economically designed and built, which includes installation of Contiguous Bored Piles (CBP) in combination of vertical grout columns behind CBP wall to seal the gap between the CB Piles, installation of inclined ground anchors, construction of capping beam, stressing and locking of installed ground anchors. A detailed monitoring program was in place and continuous monitoring of wall deflections is being measured at the time of publishing this paper. This paper describes the design and construction aspects adopted for foundation and retention system including the monitoring results during excavation stage of the project.INTRODUCTIONIn general, stage wise open excavation will be opted for basement construction where plenty of space is available. Due to urbanization and restricted space, open excavation became a difficult task nowadays. To overcome this problem, suitable retention system need to be selected to retain the soil mass. Selection of suitable retention system for multi-level basement excavation is a great challenge in geotechnical industry, especially when the proposed site is located within the minimum set back distance from the existing structures. This in turn calls for the advanced and innovative methods of retention system.Several earth retention systems such as Diaphragm Walls, Secant Pile Walls, CBP system, Soldier Pile Walls, Sheet Piles, Deep Soil Mixing (block walls) etc. with or without inclined ground anchors are being practiced worldwide. Considering the advantages and disadvantages of the available retention systems in practice locally and keeping in view of present site conditions, the developer has chosen CBP system with inclined ground anchors so that all foundation related works of the project will be handled in a single package. The main advantage of such packaged solution will ease the coordination between design consultants and foundation contractor resulting in faster construction of foundation works which in turn reflect savings in time and cost."

Jan 1, 2015

By Manqiu Xu, Songtao Yang, Zongfu Dai, Robert Pelton, Carla Abarca

"Fine mineral particles are difficult to process by flotation because of their low collision efficiencies with air bubbles. Decreasing air bubble size and increasing mineral particle size through flocculation are the most common approaches to improve flotation efficiencies. Herein we demonstrate that nanoparticle flotation collectors can be engineered to selectively induce fine mineral particle aggregation and promote flotation.Flotation tests of clean fine pentlandite (Pn) and Clarabelle Mill (CBM) rock tails and pyrrhotite (Po) tails that contained significant amount of fine pentlandite have confirmed that the nanoparticles can serve as both collectors and selective flocculants for fine pentlandite. In the case of clean fine Pn, the nanoparticle collectors virtually can float almost all Pn fines. Scoping tests on CBM tails have shown significant improvement in nickel recovery. Addition of nanoparticles as a flotation aid can recover 20% of the nickel contained in the rock tails and 40% of the Pn contained in the Po tails compared to current mill operations giving 0% nickel recovery from rock tails and no Pn selectivity from Po tails. Ongoing work includes attempts to identify optimal strategies for employing nanoparticles in fines flotation.IntroductionFine particle flotation is an industry-wide challenge in mineral processing. Fine mineral particles are difficult to process by flotation because fine particles have low mass and thus low momentum, which results in their low collision efficiency with air bubbles (Dai, Fornasiero, & Ralston, 1999). In addition, fine particles have not only high specific surface area but also more active surface area, which causes high reagent consumption. For base metal sulphide minerals, particle size between 10 µm and 70 µm is a typical range where high flotation recoveries can be achieved with conventional flotation chemistry and acceptable cell residence times. For pentlandite, the flotation recoveries decrease sharply when the particle size is below 10 µm and hydraulic entrainment is the dominant mechanism, giving recoveries proportional to water recovery and no Pn enrichment. (Jameson, Nguyen, & Ata, 2007)"

Jan 1, 2015

By O. Celep, T. Uslu, Savas

This study investigates the enrichment of a low-grade colemanite concentrate (-3 mm) using a Knelson centrifugal gravity concentrator. Due to its low boron content, the concentrate is unsaleable and has to be stored under appropriate conditions to avoid potential environmental problems. The low-grade colemanite concentrate was comminuted to size fractions of-1 mm, -0.5 mm, and -0.15 mm before treatment in the Knelson Concentrator. The effects of particle size, fluidizing water velocity, and bowl speed on the enrichment process were examined. The B2O3content of the concentrate was increased from 33.96% to a maximum of 45.52%. B2O3recovery increased with increasing bowl speed and particle size, and decreased with increasing fluidizing water velocity. The enrichment process also rejected arsenic and iron to some extent, with a maximum reduction of arsenic from 1360 g/t to 765 g/t and iron from 0.88% to0.33%.

Jan 1, 2015

By P. den Hoed, R. I. Rambiyana

"Four sulphide minerals - pyrite (FeS2), pyrrhotite (Fe1-xS), pentlandite ([Ni,Fe]9S8), and chalcopyrite (CuFeS2) - contain the base metals and most of the iron in concentrates of platinum group metals (PGMs). In the pyrometallurgical processing of PGM concentrates these sulphides form a matte during smelting, and iron and sulphur are removed from the matte during the converting process. This paper discusses the roasting of Merensky concentrate in air before smelting, with the purpose of reducing the matte load to the converter.Roasting tests were conducted in a bench-scale rotary kiln at temperatures from 350°C to 700°C. The concentrate tested contained 17.4% sulphur and consisted of 23% pyrrhotite, 16% pentlandite, 11% chalcopyrite, and 2% pyrite. The particles were fine (d50 = 22 µm), and all the sulphide particles were liberated. Roasting in air at 550°C and 650°C for 20 minutes removed respectively 60% and 70% of the sulphur. The iron in the sulphides was oxidized to Fe3O4 (magnetite) at temperatures below 500°C and to Fe2O3 (haematite) at temperatures above 550°C. At 700°C the bed sintered and copper oxides formed. At temperatures below 450°C oxidation was incomplete: pyrrhotite remained and only 30% of the sulphur was removed. Smelting tests were conducted to assess matte fall and the deportment of copper and nickel to matte. It was evident that roasting resulted in lower matte falls (a drop of approximately 60%) compared with matte falls from unroasted concentrate. The iron and sulphur levels in the matte were reduced to below 3.5% and 22% respectively.This paper also briefly describes the mechanisms by which pyrrhotite, chalcopyrite, and pentlandite are oxidized during roasting. For chalcopyrite, the mechanism proceeds through an intermediate solid solution phase, which extends from Cu1.02Fe1.04S2 to Cu2.04Fe0.72S2 to a copper-rich solid solution of bornite (Cu4Fe1.4S4-Cu2S). The oxidation of pentlandite proceeds through a monosulphide solid solution (Ni0.39Fe0.53S-Ni0.74Fe0.15S) to a solid solution of heazlewoodite ([Ni,Fe]3}xS2). These mechanisms are explored in relation to chemical thermodynamics and microstructures."

Jan 1, 2015

By S. J. Lötter, J. T. Nel, W. Purcell

"SynopsisGlow discharge optical emission spectroscopy (GD-OES) is an analytical technique used in the analysis of solid, conducting materials. Though primarily of interest as a depth profiling technique on samples with varying layers of both conducting and non-conducting materials, it is also capable of rapid bulk analysis of homogenous, solid samples. GD-OES combines the advantages of ICP-OES (wide detection range, speed, and lack of interferences) with the solid sampling of XRF techniques. This allows the analyst to not only quantify the elemental composition of a sample, but to evaluate it in terms of homogeneity with depth, a field in which auger electron spectroscopy (AES) and secondary ion mass spectrometry (SIMS) have traditionally been the primary techniques. Although GD-OES does not replace these useful techniques, it does offer various advantages over them, making it an excellent complementary analytical tool.In GD-OES analysis, a low-pressure glow discharge plasma is generated with the sample material acting as a cathode, accelerating the cations in the plasma towards the sample surface. This bombardment causes the sample material to ‘sputter’, essentially knocking free atoms or molecules of analyte material, which are then drawn into the plasma where they are excited. The light emitted by this excitation is then diffracted to separate the wavelengths emitted by the specific elements and detected by a spectrophotometer. The intensity of the signal is directly proportional to the quantity of analyte element present in the sample, allowing simple calibration and quantitative determination of the elements. Technological improvements made in the past twenty years or so have significantly increased the usefulness of GD-OES for surface analysis. Faster plasma stabilization and start-up allow for the quantification of surface layers as thin as 1 nm. Sputter rates have been accurately measured for many common materials, allowing them to be built into software libraries in the instrument’s control software. This dramatically expands the usefulness of this software and the ease of performing analyses.GD-OES analysis of nuclear materials allows for the rapid determination of the elemental composition without the requirement of initial dissolution. The thickness of any corrosion layers on nuclear materials can also be determined."

Jan 1, 2015

By K. G. Matshika

"Nickel electroplating from sulphate electrolyte is an efficient process. To achieve good quality cathode, however, parameters such as temperature, nickel concentration, pH, electrode alignment, electrolyte flowrate and additives must be optimised, because these affect cathode morphology and current efficiency. It is also important to control the clarity and contamination of the electrolyte, as these have a considerable effect on the quality of the cathode surface. Delamination, surface roughness and pitting on the cathode surface can be caused by incorporation of solid particles or impurities form the electrolyte. Three types of solid particles were investigated: cuprous selenide, barium/ lead sulphate (co-precipitated) and ferric hydroxide. These particles can be found in industrial electrolytes, especially if there is a problem with the filtration step or contaminants have reacted to form insoluble particles. A rotating cylinder Hull cell was used for this investigation. It is found that the effects on morphology depend on the type of solid, concentration, particle size, current density and electrolyte flowrate. Cuprous selenide is the most detrimental, especially at higher concentrations, higher current densities and low flowrates. Increased rotational speed offers a uniform sheet of fine gas bubbles across the surface of the cathode which reduces the thickness of the diffusion layer of liquid adjacent to the cathode and increases the rate of mass transport to the cathode from the bulk liquid. INTRODUCTION Electrowinning operations are important when recovering associated by-products of nickel production, such as copper, cobalt and zinc, from pressure leach solutions. The base metal mineralization associated with platinum-group metal (PGM) production yields significant quantities of copper and nickel as by-products. The base metal refinery is vital in PGM processing as it increases the PGM concentration from mined ore by a factor of 400, in comparison with concentration by flotation (20 times) and pyrometallurgical processing (20 times) (Crundwell et al., 2011). Anglo American Platinum Ltd produces nickel cathodes by electrowinning and purges sodium from the base metal refinery in the form of sodium sulphate at the Rustenburg Base Metals Refinery (RBMR), as shown in Figure 1. This is the Sherritt acid leach process used in most base metal refineries to ensure no loss of PGMs and shows the subsequent purification steps for copper and nickel. Other refineries in Zimbabwe, such as Bindura and Empress, use the Outokumpu leach process (Pavlides, 2001)."

Jan 1, 2015

By A. G. Doll

"Multi-stage crushing plants, including high pressure grinding roll (HPGR) circuits, require more materials handling equipment than SAG and ball mill circuits. A temptation exists to neglect the energy consumption of materials handling when doing desktop comparisons of HPGR versus SAG milling circuits because estimation of materials handling power requirements can demand significant general arrangement drafting that is not available when performing preliminary scoping or ""desktop"" studies.The largest component of materials handling power is consumed by conveyors when lifting material between stages of crushing, meaning that a simple potential energy model can be used to evaluate conveying specific power consumption. This potential energy consumption for conveying can then be factored to provide overall materials handling specific energy consumption suitable for desktop studies of HPGR and other multi-stage crushing circuits. The technique also permits different crushing circuit flowsheets to be evaluated, at a preliminary level, by simply counting the number of times a conveyor must ""lift"" ore from ground level up to a bin or other equipment mounted up high.INTRODUCTIONTrade-off studies of multi-stage crushing plants involving high pressure grinding rolls (HPGRs) versus semi-autogenous grinding (SAG) mill circuits typically focus on the operating cost savings of crushing plants versus the capital cost savings of SAG mill plants. Very high-level (or desktop) studies of this type do not include a great deal of engineering work, meaning that the details (and costs) of ancillary systems are often not captured when doing the comparisons. The more complicated nature of the multistage crushing plants means there are more ancillary systems than a SAG mill plant, risking a skewed analysis of the operating costs, benefiting the HPGR option.One of the ancillary systems that may be overlooked is the more complicated conveying systems inherent in multi-stage crushing plants. Due to the mass of material being handled in large mining operations, these conveyors become a significant portion of the operating costs. A simple way to estimate the specific energy consumption of conveying systems without requiring engineering drawings and a site layout would allow high-level studies to capture this operating cost."

Jan 1, 2015

By S. Kelebek, C. Gibson, M. Aghamirian

"Global niobium production is currently dominated by three operating mines, Araxá and Catalão in Brazil, and Niobec in Canada. Although over 90% of the world’s niobium is mined and processed in Brazil, a number of high grade niobium deposits exist worldwide. The advancement of these deposits depends largely on the development of operable beneficiation flowsheets. Pyrochlore, the primary niobium bearing mineral, is typically upgraded by flotation with amine collectors at acidic pH, and often includes a reverse gangue flotation stage prior to pyrochlore flotation. Challenges related to the flotation of niobium ores include complicated flowsheets which require pulp pH adjustment from alkaline to highly acidic conditions, difficulty recovering fine grained pyrochlore, producing concentrates that meet specifications, and minimizing the losses to reverse pre-flotation gangue concentrates, slimes, and tailings. The specific flotation challenges associated with different niobium ores are discussed. In some cases, the use of alternate flotation reagents and flowsheets are required to achieve reasonable gangue rejection and maximize recovery, particularly of very fine pyrochlore grains.INTRODUCTIONNiobium is a transition metal primarily found in the oxide minerals pyrochlore (Ca,Na)2Nb2O6(OH,F) and columbite (Fe,Mn)Nb2O6. After physical upgrading, pyrochlore is converted into a ferroniobium product which is largely used in the production of high strength low-alloy steel. Today’s niobium market is dominated by the Companhia Brasileira de Metalurgia e Mineração, whose Araxá operation produces over 80% of the world’s niobium. There are two smaller operations, Niobec in Canada, and Catalão in Brazil, which account for the remaining production (U.S. Geological Survey, 2013).A number of niobium deposits exist globally, their advancement dependent on the development of operable beneficiation flowsheets. Pyrochlore, the predominant niobium bearing mineral, is commonly upgraded by flotation with cationic amine collectors in acidic pulp conditions. However, many niobium ores do not respond to this typical flowsheet because they are either highly refractory or niobium deports to other minerals such as columbite, ilmenite and rutile, which are not as responsive to this reagent scheme (Bulatovic, 2010). Other reagents, such as alkyl hydroxamic acid, have been studied extensively and while promising, are not without their own inherent issues including lack of selectivity over iron oxide minerals, and high addition rates (Ni & Liu, 2012; Gorlovskii, Eropkiny, Koval, Strelstin, Khobotova, & Shtchukina, 1968; Gorlovskii, Khobatova, & Shtchukina, 1966).It is the aim of this report to highlight the challenges associated with the physical upgrading of niobium minerals through examination of the flotation behavior of two different ore systems. In order to overcome these challenges it is important to thoroughly understand the limitations of typical flowsheets/reagent schemes currently employed in operation."

Jan 1, 2015

By A. Patri, D. S. Nimaje

Wireless sensor networks and wireless communication systems have become indispensable in underground mines. Wireless sensor networks are being used for better real-time data acquisition from ground monitoring devices, gas sensors, and mining equipment, whereas wireless communication systems are needed for locating and communicating with workers. Conventional methods like wireline communication have proved to be ineffective in the event of mine hazards such as roof falls, fires etc. Before implementation of any wireless system, the variable path loss indices for different workplaces should be determined. This helps in better signal reception and sensor node localization, and also improves the method by which miners carrying the wireless devices are tracked. This paper proposes a novel method for determining the parameters of a suitable radio propagation model, which is illustrated with the results of a practical experiment carried out in an underground coal mine in southern India. The path loss indices, along with other essential parameters for accurate localization, have been determined using the XBee modules and ZigBee protocol at 2.4 GHz frequency.

Jan 1, 2015

By Dean Lindsay, Inna Dymov, Mariam Melashvili, Jim Smolik, Chris Fleming

"This paper presents results from the proposed cyanide detoxification circuit for Seabridge Gold’s KSM project. The proposed flowsheet uses a two-stage CCD (counter-current washing/decantation) circuit to transfer the flotation products CIL(carbon-in-leach) barren solution to a SART process (Sulphidization, Acidification, Recycling of precipitate, and Thickening of precipitate) to recover copper and free cyanide from the copper cyanide complex. The SART barren solution is then treated by an AVR process (Acidification, Volatilization of HCN gas, and Reneutralization) to concentrate the cyanide that was liberated in SART for recycling to leach. The washed CIL barren pulp (the CCD underflow) is treated with the SO2 /Air process to oxidize residual cyanide.The goal of this complex flowsheet is to recover and recycle as much cyanide as possible and produce a final effluent containing less than 0.5 mg/L copper and CNWAD at the lowest cost. The SO2/air process is capable of achieving low residual CNWAD (<1 mg/L) in the final product, but the persistence of aqueous copper above the target level has been detected in the treated solution in some instances. The expected high concentrations of copper and cyanide in the KSM CIL barren ( >200 mg/L each species) and the difficulties this might create for the SO2 /air process was a key factor in the decision to consider including the washing, SART and AVR steps in the project flowsheet.The target concentrations of CNWAD and Cu were readily achieved after aging the SO2 /air treated pulp, as would occur in a tailing or polishing pond. This suggests no additional treatment will be required to sustain low levels of these species in the final plant effluent."

Jan 1, 2015

By M. Mikolás, L. Pomothy, K. Špakovská, M. Vanek

"Enterprises are faced with increasing economic competition and managers are obliged to look for methods that will ensure a competitive edge in their companies’ markets. These methods include managerial concepts that employ common sense and are low cost. One of these approaches is the KAIZEN methodology, of Japanese origin; kaizen means continuous improvement (CI) when translated. A mining company, despite some of its idiosyncrasies, is just an enterprise where the application of KAIZEN can be advantageous. The company investigated, OKD, is the only domestic producer of hard coal in the Czech Republic and operates four mines that produce 11 Mt of coal annually. The OKD management team opted for KAIZEN as their principal method of continuous quality improvement. This paper suggests ideas for development and application of processes that provide for continuing improvement of production and management. The authors have taken the OKD Company to be their benchmark in the field of mining activities. The paper also focuses on possible difficulties accompanying the introduction of CI sustainable methods. Three years of application of CI methods has increased the income of the company by almost US$38.1 million, which provides a strong argument for continued application."

Jan 1, 2015

By C. G. Anderson

"As demand for copper and precious metals continues to grow, the need to develop new supplies increases. In the imminent future, the new large global ore bodies are more complex, with enargite, tennantite and tetrahedrite as the abundant contained mineralogy. These resources also contain considerable gold and silver. As such, direct treatment by conventional smelting is prohibitive due to the potential environmental impacts and copper quality issues. This paper outlines the effective application and associated economics of industrially proven Alkaline Sulfide Leaching (ASL) hydrometallurgical technology to an enargite concentrate for selective removal and recovery of arsenic, antimony, gold and silver, leaving a clean high-grade copper concentrate. The fundamentals behind the technology are described, along with those of hydrometallurgical arsenic fixation and stabilization. Flowsheets and unit operations are elucidated, along with an in-depth economic analysis. INTRODUCTION Hydrometallurgical methods can be employed for treatment of gold-, arsenic-, antimony-, tin- and mercury-containing materials, concentrates and ores, as well as complex ones containing any number of metals (Conner & Anderson, 2013). The alkaline sulfide leaching (ASL) system is one of these. It is essentially a mixture of sodium sulfide and sodium hydroxide with other meta-stable alkaline species as required. This is a unique hydrometallurgical system beause it is a very selective lixiviant for the distinct leaching of tin, gold, antimony, arsenic, tellurium, silver and mercury (Anderson, 2000, 2001a, 2001b, 2002, 2003, 2012, 2014; Anderson & Twidwell, 2008a, 2008b, 2010; Anderson et al., 2004, 2005; Glazkov & Tseft, 1961; Gnatyshenko & Polyvyanni, 1961; Nadkarni et al., 1975; Nadkarni & Kusik, 1988; Polyvyanni, 1963). Worldwide, ASL technology has been employed industrially in the former CIS, China and the United States for the production of antimony (Anderson et al., 1992, 1994; Anderson & Krys, 1993; Holmes, 1944; Kaloc, 1967; Nordwick & Anderson, 1993; van Stein, 1971)."

Jan 1, 2015

By M. Pineda

Sublevel stoping (SLS) is one of the oldest and most used methods for underground mining. It relies heavily on the use of drilling and blasting techniques to remove the rock, and gravity to transport the broken rock to drawpoints located at the base of the stope, with LHDs to transport material from the drawpoints. Current SLS operations are based on the assumption of stable geometry of the stope. Thus, the stope design includes the definition of the geometry according to the orebody shape and geomechanical constraints to avoid instability, which may cause excessive dilution. Under some circumstances, dilution could enter the stope due to geotechnical instability, especially when large stope geometries are used. A review of current design and planning practices for large SLS operations indicates that no consideration is given to the material flow and the mixing that occurs after blasting. Material flow could have a large impacton the mixing of ore when grades are heterogeneous in the stope. In this paper, we discuss the influence of gravity flow on the design and planning of large sublevel stopes with and without vertical dilution, based on laboratory experiments. The outcomes of this investigation are used todevelop guidelines towards the design and planning of large SLS mines, which would complement the currently used geotechnical considerations.

Jan 1, 2015

By Fadeela Saloojee, Frank K. Crundwell

"Pressure leaching is an option for copper recovery from chalcopyrite. Leaching takes place at high temperatures and pressures in the presence of an oxidising agent. This work shows how the performance of pressure leaching circuits may be improved by optimising the configuration of autoclaves and the heat-removal method. The objective was achieved by creating an autoclave model by combining mass balances, energy balances and population balances. A base-case circuit consisting of a single autoclave was chosen, and this base case was compared with options for increasing capacity by adding more autoclaves to the circuit. These capacity–increase options include circuits in which the additional autoclave is added in parallel, in series and in series with thickening between the two autoclaves. The copper extraction, productivity and cooling requirements for these options are compared. The series circuit with inter-stage thickening has the highest extraction and productivity, however, the cooling duty in the first autoclave is high. In addition to comparing circuit options, three options for heat removal were investigated: cooling coils, quench water and flash recycling. The flash-recycle option results in the highest copper extraction, and the quench-water option gives the lowest copper extraction. INTRODUCTION Background Copper sulfide ores and concentrates containing chalcopyrite are typically hard to leach under atmospheric conditions. Smelting of these ores may also be a challenge, due to the inability of smelting processes to handle low-grade ores, as well as impurities such as arsenic and antimony. Alternative hydrometallurgical routes to treating chalcopyrite are therefore required. One option is pressure leaching, which is the focus of this work. In the pressure leaching process, sulfide ores and concentrates are leached in autoclaves at temperatures of 200 to 220°C and a pressure of about 3000 kPa (Schlesinger et al., 2011). An oxidising agent, usually oxygen, is required to oxidise sulfide ions to sulfate ions. The oxidation of sulfur to sulfate is exothermic; consequently, a method of heat removal is necessary in order to maintain the temperature in the autoclave."

Jan 1, 2015

By Brian T. Burch, Paul N. Worsey

When fired, submerged Linear Shaped Charges (LSCs) lose all effectiveness in the cutting of steel. Users in underwater applications have reported having to substantially increase the charge size to the point where brute explosive force shears rather than cuts the target.

Jan 1, 2015

By P. Tolvanen, A. Jalili

"When hard times strike in the mining cycle, sites tighten their belts and focus on lowering their unit costs. Lower unit costs can be achieved by reducing expenditure, increasing production, or by a combination of both. Faced with these requirements, site personnel must turn to areas where they can focus their effort for maximum return. For many sites, both in Canada and overseas, focusing on improving flotation cell maintenance is one of the areas where gains can be made for minimum outlay. Due to resource limitations and/or lack of equipment understanding, maintenance on many sites is only conducted on a reactive basis, once equipment fails. This paper will promote the idea that mineral processors should become more familiar with their equipment so they can ensure that the plant’s ability to produce is not compromised. By better understanding the flotation equipment and ensuring the critical components are maintained, it should be possible to unlock value in an operation. Flotation equipment inspections will be discussed in detail, including key items to look for during these inspections. Potential approaches for predicting wear and change-out time for flotation equipment components will be also demonstrated. An example of one approach used at a South African nickel site will be given.INTRODUCTIONWith metal prices decreasing and operating costs rising, concentrator operators are looking for ways to improve their bottom line. Lower unit costs can be achieved by reducing expenditure, increasing production, or by a combination of both. To meet the expectation of shareholders, mineral processors must turn to areas where they can get the best return for their efforts. One area where there is a large opportunity for improvement is in the maintenance of flotation equipment.On many mine sites, both domestically and around the world, it is quite common to see the philosophy applied of repairing equipment only once it has failed. Though some sites have advanced enterprise management systems and monitoring programs, in the authors’ experience, the decision on flotation equipment maintenance is rarely made on a proactive basis. Two common trends that are evident in the mining industry are that operations are under pressure to keep the plant producing, and maintenance managers generally have very strict budget targets to meet. In addition to these factors, few sites have a good understanding of the financial implications of running with worn or broken flotation equipment. Knights, Jullian and Jofre (2006) summarize this situation well, stating that the lack of correlation between some commonly used maintenance performance measures and key business drivers can lead to maintenance being seen only as a cost center (that needs to be reduced) rather than a tool for productivity improvement."

Jan 1, 2015