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By Vilkul Yuri

Operation of the iron ore deposits of Krivbass which has been going for 130 years, has led to serious problems related to reaching great depths in open pit and underground mining. Extraction of the huge volume of rock mass (about 9 billion m3) resulted in mass violations of the natural state of the subsurface. At present most mining operations are carried out either by combined open pit and underground mining or by open pits in areas underworked by underground mining. Now the process of movement of the strata underworked by underground mines is not completed ? work fronts of open pits approach the areas of possible funnel, terrace and fissure formation, movement continue to develop. With deepening mining operations the remaining reserves are worked off under layers of incumbent rocks of > 900 m. Significant underground voids arise from mining "blind" deposits at large depths. Heaving of voids over the mined area lasts for decades; it may result in not only unpredictable time of funnel emerging and area development of movement synclines, but also the threat of air blows in mines. Forced caving of incumbent rocks as a way of controllable backfilling is difficult to achieve for economic and technical reasons. This is much more expensive and dangerous than stoping. Furthermore, the mining practice does not know examples of enormous volume backfilling under a km thick rock layer. Currently, RMI "KNU" is conducting research and pre-design work directed at creating a specialized pilot Geoinformation System (GIS) of continuous remote monitoring of rock mass in areas of underground voids influence in Krivbass area. Implementation of this system will enable to determine promptly in time and space the areas of the beyond cut off stress condition, to inform about possible occurrence of events that may result in accidents and to improve safety of mining along with other systems of forecasting and controlling the stress condition in rock mass. RMI "KNU", with a large number of professionals from many Ukraine?s leading mining research centers, has developed a research and design work program that includes the following principal stages: 1. Analyzing geomechanical conditions of combined mining of magnetite quartzite and rich ore deposits; determining mechanisms of moving of rock strata underworked by underground mining and places of hazardous area monitoring. 2. Investigating anthropogenic factor influence on the safe state level of rock strata underworked by underground mining (laboratory studies). 3. Developing processing techniques for continuous remote monitoring of rock mass in areas of underground cavity influence. 4. Modeling the state of rock mass based on results of pilot GIS hardware measurements. 5. Developing an expert diagnostics system for rock mass state in areas underground void influence. 6. Experimental investigation and testing the GIS, its improving for efficient application. After testing the pilot GIS it is supposed to be implemented at all mining enterprises as an integral part of iron ore mining and processing technology in the Krivbass area.

Aug 1, 2013

By X. J. Guo, H. X. Li

The biological heap leaching at Zijin copper mine is currently the largest industrial application of biohydrometallurgy and one of the earliest mines to develop biological metallurgy technology in China. Since 2001, through demonstration plant and different expansion stages, it has been developed to produce 20,000 tons cathode copper annually. However the ore grade of heap leaching gradually decline year by year with copper grade average at around 0.23% now. The copper mineral is mainly the secondary copper sulfide, and symbiosis with higher content pyrite, the sulfur and iron contents is about 15 times as copper in the ore (S/Cu>15), so the acid and iron treatment of leachate, extractant and acid mine waste water is a great challenge. Combined with located in the rainy area of southern China, huge amounts of over 10 million cubic meters acid drainage need to be treated every year, so the recovery of copper from waste water as well as low grade ore is one of the characteristics of this process. This paper in general introduced the operation of heap leaching, solvent extraction, stripping and electrowinning process in Zijin mine, summarized the experience of production and technical practice, and discussed the methods of heap construction, leaching rate enhancement, third phase formation and treatment, mine effluent treatment, resource recovery and heap slope reclamation. INTRODUCTION The Zijinshan gold and copper mine is located in southeast China, Fujian province, and the typical distribution characteristics of gold and copper deposit are "upper gold and lower copper", both are large porphyry deposits. By the end of 2017, the accumulated proven reserves of available gold resources were 318 tons, copper are 2.15 million tons, and the prospective reserves of copper are more than 5 million tons. The gold mining is open pit while the copper mining is open pit and underground. The cut-off grade of gold mining is 0.15 g/t, and that of copper mining is 0.15%. A large number of very low-grade gold ores (grade <0.30 g/t) adopt a combined ore-dressing process of “partial crushing/ore washing/reprocessing/heap leaching (overlying)/ carbon slurry" independently developed, while flotation and bio-heap leaching are adopted for copper mine. Among them, copper ore with more than 0.25% copper content enters the flotation system to produce copper concentrate, while that of less than 0.25% copper content is processed by bio-heap leaching, solvent extraction and electrowinning. Zijin biological heap leaching plant is the first mine in China to achieve the scale of annual output of ten thousand tons of cathode copper since 2006. In 2008, the solvent extraction and electrowinning system were expanded, now with a daily processing capacity of 20,000 tons of ores, an annual output of 20,000 tons of cathode copper, the expanded plant is the largest biohydrometallurgy plant in China.

Jan 1, 2019

By G. D. Hibbard

New regions of material property space can be accessed by combining microstructural design at the nm-scale with architectural design at the µm- or mm-scale. In the first case, large strength increases associated with grain size reduction to below 50 nm have driven extensive research efforts into the development of nanocrystalline materials [e.g. 1-3]. For many potential structural applications, however, the density of a nanocrystalline material is just as important as its strength. In fact, reducing the density is more important than increasing the strength for certain weight specific materials performance indices, and it is especially critical for applying structural nanomaterials in the aerospace and automotive sectors. We have developed a new type of structural nanomaterial wherein the effective density of the parent metal is reduced by more than an order of magnitude by incorporating an internal periodic cellular architecture of open space. In one example a low density cellular nanocrystalline material was created by electroforming nanocrystalline Ni around a rapid-prototyped acrylic photopolymer microtruss, Figure 1 [4]. Microtruss materials have periodic cellular architectures that are specifically designed to undergo stretching-dominated deformation as opposed to the bending-dominated deformation in conventional open cell metal foams [5,6]. This new cellular nanocrystalline hybrid material combined the structural efficiency of microtruss architectures with the ultra-high strength that can be achieved by grain size reduction to the nm-scale. Although it played a critical role as a cathode support during the initial stages of nanocrystalline electrodeposition, the polymer core did not contribute significantly to the inelastic buckling resistance of the composite metal/polymer struts and it may therefore be desirable to remove it post deposition by means of chemical dissolution or thermal decomposition [4]. Nanocrystalline electrodeposition can also be used to reinforce conventional cellular metals (Figure 2), creating new types of cellular composites such as metal/metal microtruss hybrids [7-10] and metal/metal foam hybrids [11,12]. For example, nanocrystalline Ni-Fe alloy sleeve thicknesses of up to 400 µm were used to reinforce an aluminum alloy microtruss that had been fabricated using a stretch-bend approach; the nanocrystalline sleeves had the effect of increasing the microtruss strength by up to a factor of twelve and the specific strength by a factor of three [8]. Alternatively, much thinner sleeves can be electroplated to produce structural coatings. In one example, a nanocrystalline Ni coating was designed to provide both corrosion protection and inelastic buckling resistance [10]. Because the ultra-high strength material was optimally located at the furthest distance from the neutral bending axis, only a thin coating of 50 µm was needed to double the inelastic buckling resistance of the 1.13 mm x 0.63 mm cross-section plain carbon steel struts [10]. Overall, the mechanical performance increase that can be obtained through electrodeposition depends on a complex interaction between the starting cellular architecture and the nanocrystalline sleeves.

Jan 1, 2010

By J. H. Starkey, P. A. Scinto

"In Canada and other countries, legislation dealing with accurate factual disclosure for mining projects is being expanded to include more metallurgical design information. Grinding characteristics, metallurgical recovery of contained metal and environmental hazards associated with tailings disposal will be part of the Canadian package. This paper will examine only that portion of the plant design standard practice that relates to grinding – in SAG mills and ball mills that grind ore from as received crushed material to final ground product that is fed to downstream plant recovery processes. Since grinding plants represent a significant portion of the overall concentrator cost, accountability in grinding design should no longer be left to chance, opinion, or hidden proprietary information. This is the 21st century. It is reasonable to expect that a factual basis will be needed so that a Qualified Person (QP) reviewing the grinding mill design can understand it. That will also involve how many and what quality of samples are taken to define the characteristics of the deposit, the accuracy of the tests used to measure ore hardness and the method by which the dimensions and power of the mills are determined. If these criteria are not met, the QP may have no choice but to call attention to the lack of convincing information regarding the grinding mill design for the project. At the same time clients are expected to demand getting the correct design information at a cheaper price, results in a more timely way, and accurate design to the point where the risk of tonnage shortfalls are reduced or eliminated.Tables will be presented to show how this can be done to current standards using existing technologies. At the same time it is acknowledged that current standards are ill-defined and that to achieve consensus among those involved in SAG mill design work, it may require that a consortium funded project be set up to establish what tests and methods for grinding design are suitable to meet the new standards for metallurgical disclosure. By running parallel design methods on several major projects over a period of 5 years, the process of how to design grinding circuits will be better understood because the projects will start at the design stage, include procurement of the mills and will end with a benchmark test on the plant.Selecting the samples and the best schedule for doing test work on core samples submitted and mill design, will be discussed in this paper. The issue of completing the grinding design in time for early procurement decisions is on the critical path for most major projects. Discussions between mining engineers, geologists and process engineers are needed early in the development cycle, at the scoping study level (SS), preliminary engineering assessment (PEA), prefeasibility study (PFS), and at the feasibility study stage (FS), to ensure that proper planning for design sample selection is done. Using adequate sampling and testing methods in a timely way allows early procurement of the grinding mills and minimizes project risk."

Jan 1, 2012

By Virginia Lawson, Adam Price

The rougher flotation stage is used to reject liberated gangue at a coarse size. Base and precious metals due to their high specific gravity tend to recycle in the grinding circuit and report to the flotation circuit at a finer size than the gangue from which it is being separated. Generally liberation levels of greater than 50% are achieved in roughing and thus a portion of the feed can be immediately upgraded to final concentrate. The proportion increasing as the amount of liberated mineral in the feed increases. Scalping out concentrate from the rougher can reduce the size of the subsequent regrind and cleaner circuits. Examples of alternative process design for circuits demonstrates that you don’t need large volumes to effectively recover all of the minerals – you need smart flotation. Several case studies will be presented where circuits can be significantly reduced in size by designing smarter – not larger. This paper describes case studies to test the amenability of the Jameson Cell to produce final concentrate from rougher feed. Metallurgical performance and plant design implications are discussed. Froth flotation has been used extensively for mineral concentration for over 100 years. Although the flotation cell design has changed, it has generally simply increased in capacity rather than increase in efficiency. Cell capacities of over 600 m3 are now being used in rougher flotation particularly for the lower grade ores where the amount of surface area for flotation is deemed less significant than the required volume. The plant design volume in rougher flotation comes from laboratory batch flotation tests and/or pilot plants. The volume required for the plant is calculated based on the laboratory residence time multiplied by a scale- up factor. Examples of laboratory to plant scale-up design are given in Dunne et al. (2010 part 2). The factors are usually between 2 and 3. A laboratory residence time of 10 minutes would equate to a plant design residence time of 25 minutes using a 2.5 factor. The volume of the required flotation cells can then be calculated using the design tonnage and volumetric throughput. This scale-up factor takes into account the differences in performance between batch flotation efficiency and full scale plant. The differences are likely related to inefficiencies of bubble particle contact in large cells as well as froth recovery, which is essentially complete in a batch cell. Flotation cells, such as the Jameson Cells, are smarter designs that improve bubble/particle collection, bubble/pulp separation efficiency and operate with high froth recovery rather than simply getting bigger. The Jameson Cell was designed on the principal of improving the collection efficiency of fine particles by separating the contact zone from the separation zone and having the bubbles generated by the plunging jet shearing and entraining air. This means the method of bubble generation is not impacted by the wear of mechanical equipment as it the case in other flotation devices. Jameson Cells make bubbles less than 600 micron in size under most operating conditions (Ding, 2016).

Jan 1, 2019

By Xian Jian Guo, Lei Li, Hua Wang

The distribution and behavior of As, Pb and Zn elements in the copper smelting process were investigated using thermodynamics. The differences between thermodynamics evaluation and practical smelting processes were analyzed. The main factors affecting the distribution of the elements are oxygen enrichment in blowing, copper matte grade and smelting temperature. In general, the distribution ratio of As, Pb and Zn to off gas increases with the smelting temperature and decreases with oxygen enrichment in blowing. Meanwhile, their distributions in matte decreases and that in slag increases with the increase of copper matte grade. There is some gap in distributions between thermodynamics evaluation and practical smelting processes. INTRODUCTION The copper concentrates usually contain a number of minor elements, the control of which in copper smelting processes is often a key to the quality of the anode copper produced and may also have a bearing on the overall economics of the process (Yang et al., 2017). Knowledge of the thermodynamic properties of the minor elements in slag is very important in the control of the slagging reactions (Takeda Y, Ishiwata & Yazawa, 2007). In addition, the environmental contamination caused by the minor elements associated with copper-bearing ores, such as As, Pb and Zn, has become more and more important (Wang, 2004). Arsenic, lead and antimony, which are all carcinogenic metalloid elements, are usually contained in iron-rich wastes, and theirs emissions are considered to be one of the foremost environmental concerns (Wang, 2004). It is important to elucidate the elements distribution of them in the reducing-matte smelting process. The distributions of As, Pb and Zn elements in the copper smelting process are dependent on many factors, including the temperature, raw material composition, reduction atmosphere, concentration level, and slag chemistry etc. (Chen L Z, Xu X, 2016). Generally, these minor elements mainly distribute in phases of gas, copper matte and slag in the smelting process (Shu, Wang, & Qu, 2016). The arsenic exists mainly in the form of sulfo-arsenides in the copper concentrate, and it will decompose and be oxidized in the smelting process (Chen & Lu, 2016). Part of the products, including AsS, AsO, As2 and As2O3, evaporates into the gaseous state, and some other part enters into slag phase in the form of As2O3 and As2O5 (Wang, Guo, & Tian, 2017). Meanwhile, the As2O3 in the slag phase can also be reduced to metallic As and combine with Cu, S, Fe forming Cu3As2, As2S3 and Fe3As2, and then enter into copper matte. During the copper smelting process, about 50% of zinc enters into slag phase in the form of ZnO and 45% exists in the copper matte in the form of ZnS. Little volatizes into the gas phase. Most lead exists in forms of PbS and PbO in the copper concentrate, and it is mainly distributed in copper matte in the smelting process (Zheng & Lin, 2015).

Jan 1, 2019

By F. Q. Su

Underground Coal Gasification (UCG) is a promising environmentally acceptable clean-coal technology. A typical UCG system has two wells drilled from the surface into a coal seam with some separation of injection from production. The wells are connected underground by various linking techniques. After creating a combustion reactor in an underground coal seam, air and oxygen flow through the injection well. Heat energy and gases are collected from the production well. As gasification progresses, the combustion reactor is moved along the linking hole. Using this process, the fracturing activity inside the coal seam serves an important role for enlargement of the gasification zone because the surface area oxidization increases continuously by coal cracking. For effective coal gasification, fracturing activity must be controlled. Moreover, excess fractures inside the coal seam and surrounding rock can induce gas leakage, underground water contamination, subsidence, etc. Therefore, monitoring and control of fracturing activity in underground areas constitute key technologies for efficient and safe UCG. To monitor fracturing activity, we conducted laboratory and small-scale field experiments of UCG using acoustic emission (AE) measurements. UCG model tests were conducted for linking and coaxial types UCG models. Coaxial UCG systems are compact and are anticipated for application in underground coal seams having a complex geological structure. The UCG model consists of a coal block molded in a steel canister (20 liter capacity) with heat-resistant mortar, with a linking hole or a coaxial-hole passing through the coal block. During coal combustion, coal temperatures, production gas contents, and AE activity were monitored under the control of injection-air and oxygen flow rates. Results show that the combustion propagated inside the specimen along the linking hole. Many AE events were generated during coal combustion. The AE activity was related closely to temperature changes inside the coal block. This AE generation apparently results from crack initiation and extension around the coal combustion area under the influence of thermal stress. However, comparing the results from UCG models of linking and coaxial types, the gasification of the linking type model showed an average calorific value as high as 11.3 MJ/m3. The coaxial type yielded an average calorific value of produced gas as only 5.79 MJ/m3. In the coaxial type of UCG model, the combustion region was limited to the area around the coaxial hole. The AE activity was lower than in the linking type model. For small-scale UCG field experiments, linking and coaxial types were tested at an open pit coal mine in Hokkaido, Japan. Using a linking type UCG, 58.2 kg coal was processed in 39 h, with the calorific value estimated as 11 MJ/m3. However, the coaxial type UCG processed 7.0 kg coal in 9 h, yielding an estimated calorific value of 6.0 MJ/m3. These tendencies resemble those of results of laboratory UCG experiments. Results of these experiments demonstrate that AE monitoring is a useful tool to evaluate combustion reactors for efficient and safe UCG systems. To develop a coaxial UCG system, results show that fracture control around combustion reactors is indispensable.

Aug 1, 2013

By Adrianus (Adrian) Halim, Tomas Bolsöy, Stina Klemo

"This paper describes the ventilation system used in underground hard rock mines in Sweden and Finland, which is very different than the system used in other major mining countries such as Australia, Canada, South Africa, and USA. The system utilizes auxiliary fans with Variable Speed Drive to distribute primary airflow to working levels instead of regulators that are commonly used in other countries. Descriptions about Ventilation on Demand in this system and ventilation regulations in Sweden and Finland are presented in this paper. This paper also quantifies a comparison between the Nordic and the system used in other countries. KEYWORDS Auxiliary fan, Regulation, Ventilation on demand (VOD), Variable speed drive (VSD) INTRODUCTION The primary objective of an underground mine ventilation system is to provide sufficient airflow to dilute atmospheric contaminants that are produced by mining activities. Typically, a mine ventilation system consists of one to four primary fans, regulators to control primary airflow distribution to levels or panels, and auxiliary fans that distribute airflow to dead-end workings via ducts. Depending on circuit geometry, underground booster fans are sometimes used. This system is widely employed in many countries such as Australia, Canada, USA, United Kingdom, and South Africa. However, mines in Nordic countries such as in Sweden and Finland employ a different system, in which auxiliary fans instead of regulators are used to control primary airflow distribution, as well as distributing airflow to dead-end workings. This paper describes the general layout of Nordic ventilation system, its VOD and heating system, its comparison with the system that is widely used in other countries which is referred as “Non-Nordic ventilation system” in this paper, and the current ventilation regulations in Sweden and Finland. The comparison was quantified by doing simulations using Ventsim Design 5 software. DESCRIPTION OF NORDIC VENTILATION SYSTEM General System Layout Figure 1 shows a schematic of typical ventilation system used in Swedish and Finnish mines. The schematic was drawn by Franzen, Myran, Larsson, and Rustan from Stiftelsen bergteknik forskning (Swedish Rock Engineering Research Foundation) and the English translation is written next to the Swedish terminologies. The system is generally a push-pull system. None of Swedish and Finnish mines use exhaust system. As shown in Figure 1, the primary airflow is distributed to levels by auxiliary fans that are bolted to a bulkhead located in the access to intake raise/shaft. These intake fans are attached to ventilation ducts that distribute airflow to all working faces. The exhaust auxiliary fans, which are not attached to any ducts and also are bolted to a bulkhead located in the access to exhaust raise/shaft, then flow the exhaust airflow into the exhaust shaft. A small quantity of this exhaust air is directed into the ramp, which acts as another main exhaust airway. The deepest level is also equipped with booster fan(s) that draw air from the intake shaft/raise and direct it into the ramp to provide fresh air in the ramp. There is no free split in this system. Figure 2 shows a schematic of a level ventilation layout in the Nordic system. Figure 3 shows a photo of these auxiliary fans."

Jan 1, 2019

By C. B. Ndimande, C. Ntsele, M. Lisso, L. Arlow

"In the processing of UG2 platinum ore, stirred milling technologies are dominating regrind applications that require grinding the particles to P80s’ in the region of 45 µm. For this application stirred mills are 30–40% more energy efficient and require a smaller foot print compared to traditional tumbling mills. In this work a study was carried out using a batch pin mill and three scales of Stirred Media Detritors. A laboratory scale pin mill, a pilot scale Stirred Media Detritor with an installed power of 18.5 KW and two industrial scale Stirred Media Detritors in continuous operation, with an installed power of 355 KW and 1,100 KW were used. The industrial scale SMDs were in the secondary grinding section of a mill floatmill float (MF2) UG2 processing plants. The tests were focused on assessing the effect of feed rate and percent solids on power consumption and product fineness at lab scale, floatation response of the SMD product at pilot scale and on an industrial scale, the effect of feed rate on energy consumption and product fineness. The measurements obtained were feed flow rates, mill power and from the samples cut, particle size distribution of both feed and product. Specific energy was observed to increase with increasing percent solids over the 30wt% to 80wt% studied. Floatation response tests of the pilot scale SMD products show improvement in grade. Specific energy and reduction rations are observed to decrease with an increase in feed rate.INTRODUCTIONThe purpose of comminution in most mineral processing applications is to liberate the valuable minerals from gangue so that they can be recovered in subsequent separation stages. Most precious and base metal plants treat low grade fine-grained ores which require a higher degree of size reduction to yield the necessary degree of liberation (Partyka & Yan, 2007). Many authors have shown that the relationship between product size and grinding energy consumption is exponential, the finer the product from the comminution device, the higher the energy consumption (Gao & Weller 1993; Napier-Munn, Morell, Morrison & Kojovic, 1996; Jankovic, Valery & La Rosa, 2003). The suggested reason for this kind of relationship is that smaller particles have fewer flaws and higher strength than larger particles. Particles with fewer flaws and higher strength require higher energy for breakage to occur. One of the challenges most plants face is that of efficient energy utilisation in the liberation of the mineral from these complex ores. Conventional tumbling mills are not energy efficient in fine and ultra-fine grinding applications because some of the energy is spent on moving the charge and does not go directly into breakage. Stirred mills have been found to be about 30% to 40% more efficient and require a smaller foot print compared to traditional tumbling mills for regrind applications that require product size with P80s’ in the sub 45 µm regions (Radziszewski & Allen, 2013). Stirred mills are able to produce fine product because they can achieve a high number of stress events per unit time and volume compared to tumbling mills (Kwade, 1999a). To minimise energy usage without compromising product quality, the mining industry has adopted stirred milling technologies for fine and ultra-fine grinding (Radziszewski & Allen 2014; Kwade, 1999b). The south Africa platinum group metal (PGM) industry alone had over 40 stirred media mills installed by 2011 (Rule, 2011). Of that 40, 13 were in the ultra-fine grinding (UFG) application and the rest in main stream inert grinding (MIG)."

Jan 1, 2016

By M. M. Papic, G. M. Kimber

"Hat Creek coal was treated by the liquid solvent extraction process in a continuous plant at the U.K. National Coal Board Coal Research Establishment.Extraction levels of over 80 per cent dry-mineral-matter-free (dmmf) coal were achieved.Total light oil yield was around 47 per cent of the dmmf coal. The light oil was shown to be a good feedstock for secondary refining into gasoline, jet fuel and diesel fuel.In spite of the high mineral matter content of the coal the solid residue was successfully removed from the coal extract by filtration.IntroductionBritish Columbia Hydro (BCH) and the National Coal Board of Great Britain (NCB) undertook a joint project to test the Hat Creek coal from British Columbia in the NCB&apos;s coal liquefaction process designed to produce transport fuels and feedstocks for the chemical industry by the liquid solvent extraction of coal(1), In this process most of the coal is dissolved in an aromatic solvent and the residual solids consisting of mineral matter and undissolved coal are removed by filtration. The resultant coal extract solution is fed to a hydrocracker for conversion to light liquid products and a fraction suitable for recycling as solvent. The light liquid products can be converted to gasoline and diesel fuel in a secondary refining stage.The dissolution of the coal is effected by heat-treating a slurry at a temperature in the region of 400°C which causes some bonds in the coal to rupture with the formation of radicals(2,3). These radicals are stabilized by hydrogen transfer from the hydroaromatic groups in the solvent, resulting in a reduction in the molecular weight of the coal. In the absence of donor groups the radicals combine to form an insoluble high molecular weight product(4). From studies of the kinetics of these reactions it has been shown that the formation and stabilization of the radicals is rapid at temperatures around 400°C(2,3). The dissolution of the depolymerizing coal is, however, relatively slow and the need to ensure complete dissolution in order to achieve a high extraction restricts the selection of digestion time(5). Further, it is important that digests do not contain depolymerizing coal as it may re-form under the applied stress in the filter cake inhibiting flow through the bed and resulting in low filtration rates(6).Thus, in practice relatively long digestion times are generally required in order to ensure complete dissolution of the depolymerizing coal, while it is also important to avoid depletion of the hydrogen donor groups in the solvent which would lower the rate of hydrogen donation to the coal and thus reduce the extraction yield and adversely affect the filtration(7). Clearly it is desirable to optimize the digestion time and temperature with regard to the extraction yield and filtration properties.The precise control of the distribution of products from the hydro-cracker is important for the composition of the recycle process solvent. The principal functions of the solvent are to provide a source of donatable hydrogen and a vehicle for the dissolution of the depolymerizing coal. Hydrogen transfer from the solvent, achieved by the conversion of hydroaromatic to aromatic groups, stabilizes the radicals formed by thermal cleavage of some of the bonds in the coal resulting in a progressive reduction in the size of the coal molecules(8,9) and dissolution in the solvent. The combination of radicals would result in the formation of insoluble material and a reduced extraction yield of the coal(4). Thus to achieve a high extraction yield the solvent should contain sufficient hydrogen donor groups to prevent extensive radical combination. The composition of the solvent should also be controlled to allow dissolution of the depolymerizing coal and to prevent precipitation of this material during the subsequent processing of the coal digest and extract solution. The formation of a precipitate not only reduces the extraction yield of the coal and adve"

Jan 1, 1985

By Mohammed Rachek

Résumé : la mine d?Elhammam, opération du groupe minier Marocain Managem, est située au sud de la ville de Meknes à 63 km, elle exploite un gisement de fluorine depuis 1974. La minéralisation se présente sous forme de deux filons principaux orientés N45° avec un pendage variant de 65 à 90° NW, entre ces deux filons on note la présence d?une succession de filons satellites sécants. La méthode d?exploitation pratiquée depuis le démarrage de la mine est celle des chambres magasins, elle assure une cadence d?extraction annuelle de 250000 tonnes. Cette méthode favorisée par une très bonne compétence mécanique des épontes de la minéralisation a offert à l?exploitation une rentabilité très satisfaisante grâce à son faible coût mine. Les travaux de recherche géologique menés en parallèle à l?exploitation ont permis la découverte d?un nouveau secteur ?K-Est Gouaïda? (KEG) qui fait l?extension nord-est du gisement principal, la minéralisation de ce secteur se présente sous forme de deux filons très étendus latéralement (900 m d?extension) et présentant une puissance minéralisée irrégulière et très faible 1.2 m en moyenne, les épontes de la minéralisation sont constituées de schistes de qualité mécanique très faible. Vers la fin des années 1990, l?exploitation s?est orientée vers le développement minier du secteur KEG. Ce développement consiste à faire le choix adéquat de la méthode et de l?infrastructure minière pour l?exploitation de ce nouveau secteur, KEG. La méthode des chambres magasins, pratiquée dans le gisement principal avec un faible coût mine, n?est pas praticable dans le nouveau secteur à cause de la qualité de ses terrains, l?irrégularité de la morphologie de sa minéralisation et de la faible puissance des ses filons. Seule la méthode des tranches montantes remblayées (TMR) peut être applicable techniquement vu les spécifications suscitées de la nouvelle minéralisation, la TMR doit être appliquée sous sa variante mécanisée pour faire face au niveau exigé de la production. Après le pilotage de l?exploitation par TMR mécanisée dans le secteur KEG, il a été confirmé que cette méthode donne une grande satisfaction en matière de stabilité des terrains, sécurité, sélectivité, productivité et récupération du gisement, néanmoins elle reste pénalisée par un coût mine trop élevé pour garder le niveau de rentabilité recherché par l?exploitation d?un produit de faible valeur marchande telle que la fluorine. Dans l?objectif de rentabiliser l?exploitation du secteur KEG qui représente en réserves l?équivalent d?une durée de vie de 4 à 5 ans et d?assurer la pérennité de l?activité, l?exploitation en collaboration avec Reminex, la filiale d?ingénierie de Managem, a adapté une nouvelle variante de la TMR mécanisée dite TMR avec plan incliné sur remblai. Cette nouvelle variante de la TMR consiste à réduire au maximum les travaux dans le stérile en reconstituant un plan incliné sur le remblai à l?intérieur des filons pour faire évoluer en vertical l?exploitation des tranches en montant vers la tête d?un panneau de minerai. La variante des tranches montantes remblayées mécanisées avec plan incliné sur remblai a été imaginée et appliquer en 1978 dans la mine de fluorine de Burc située au Sud-Ouest de la France (Sicard et Mailloux, 1983). Imaginée pour une cadence annuelle de 30000 tonnes et une mécanisation très modérée cette variante a été adaptée par l?exploitation de la mine d?Elhammam en 2002 pour un haut degré de mécanisation et une cadence d?extraction annuelle de 250000 tonnes de minerai. Cette variante de la TMR est appliquée depuis le début de 2003 dans les chantiers du secteur KEG, elle présente les mêmes avantages techniques que la TMR mécanisée classique et elle a permis la réduction du volume des travaux d?infrastructure de 17.65 m à 6.25 m de travaux par 1000 tonnes de minerai. Ceci s?est traduit par une baisse du coût technique de la mine de 34% ce qui a donné un nouveau souffle à la mine pour assurer sa production et pour poursuivre la recherche géologique d?éventuelles zones à mettre en valeur.

Oct 1, 2004

This paper examines the latest advances in mine hoist motor and drives systems. The technology employed for large hoists has migrated from the traditional DC technology to AC motors with cyclo convertor drives, and more recently, advances in pulse width modulation (PWM) drives have resulted in these being applied to large hoist drives. When combined with PWM drive technology, the induction motor presents the ideal solution for hoist applications This paper examines the factors in choosing a suitable PWM drive and motor for a hoist installation. Factors under consideration include: ? hoist duty cycle, ? drive rating restrictions, ? choice of power semiconductor device,, ? low speed operation, ? maintenance ? reliability The cage induction motor and synchronous motor are compared and it is shown that the induction motor is the motor of choice for large hoists.

May 1, 2010

By L. Bichler

Utilizing thermal analysis techniques, pioneering research into the solidification mechanism of a popular magnesium alloy AZ91E cast by the Lost Foam Casting (LFC) process was carried out, Foam characteristics, section thickness and application of vacuum were seen to have profound effect on the cooling rate, thermal gradients and casting temperature profiles.

Jan 1, 2005

By K. Ghyselbrecht, R. Swennen, Y. W. Chiang, V. Cappuyns, R. M. Santos, T. Van Gerven, B. Meesschaert

Heavy metal contamination from mining and metallurgical activities of aquatic systems poses significant risk to the environment and human populations. The investigation of the distribution and immobilization mechanisms of heavy metals in sediments and soils has recently attained high priority. Different remediation methods can be employed to suppress or eliminate these contaminants. A tailored approach using both commercially available and in-house made adsorbents to rem ediate a heavy metal contamination site due to historical metallurgical activities in north Belgium is studied in this paper.

Jan 1, 2011

To date, improved thickening of clay mineral dispersions presents major technological and environmental challenge to mining and mineral processing industry worldwide. Flocculant-assisted dewatering processes currently employed during thickening are far from being efficient, requiring further development and process optimisation for dramatic improvement. In this paper, recent studies carried out to modify (i) pulp chemistry, (ii) flocculation performance and (iii) particle interactions to significantly improve the dewatering behaviour of flocculated, model and plant smectite and kaolinite clay dispersions are presented. It is shown that pulp pH-mediated hydrolyzable metal ion hydrolysis and specific adsorption in tandem with high molecular weight flocculant and shear may synergistically act to significantly enhance the dewaterability of both clay mineral pulps. Furthermore, polymeric flocculant structure type is demonstrated to show clear links with floc shear sensitivity, and hence, flocculation performance and pulp dewaterability. The implications and relevance of the findings plant clay mineral waste tailings thickening operations are drawn.

Jan 1, 2006

By K. Poorhaydari, B. M. Patchett, D. G. Ivey

In this paper, some applications of carbon extraction replicas, as samples for transmission electron microscopy (TEM), are discussed. Modern high strength microalloyed (HSMA) steels, used as structural or pipeline materials, have very small grains with substructures. Replicas used in TEM can resolve the grain boundaries and can be used for grain size measurements in cases where the small size of the grains pushes the resolution of conventional optical microscopes. The grain size variations obtained from replicas correlated very well with mechanical property variations, indicated by hardness, in the fine-grained heat-affected zone and base metal of some welded HSMA steels. Replicas are also very suitable samples for precipitate examination with analytical TEM. Advantages and limitations of the application of replicas, as well as some corrections for the estimation of the volume fraction of fine precipitates, are discussed.

Jan 1, 2005

By E. Poiate, P. A. L. P. Firme, D. Roehl, M. A. Carrion, A. M. Costa

Drilling high-pressure/high-temperature wells in salt rock formations is still a challenge for the energy industry. Creep and solubility is part of the salt rock physical-chemical behaviour, which must be considered in the design of the wells. The stress redistribution caused by drilling, in addition to thermallyactivated processes, triggers creep deformation, leading to borehole closure and the imposition of loads on the casing. The drilling fluid is designed by computer modeling in order to control the creep strain rate to a limit which allows the drilling operations and the casing setting. Depending upon the composition of the drilling fluid, chemical reactions on the salt wall may occur, especially in case of water-based fluids in highly soluble salt rocks. In those cases, cavities by dissolution (washouts) can be generated and the casing column remains uncovered by the cementation and laterally unconfined along the cavity length. This scenario is propitious to the loss of integrity of the casing due to phenomena such as plastic collapse, buckling and non-uniform loading on the casing provoked by the closure of salt formation around it. The casing should remain intact during the wellbore life-time. In this work, a section of a typical well crossing salt formation composed by halite and carnallite is considered. An idealized-cylindrical washout cavity inside a highly soluble salt layer is assumed. The purpose here is to evaluate the effect of the dimensions of the cavity on the integrity of the casing. Several typical diameters and lengths of the cavity are adopted. Numerical analyses are carried out with an in-house finite-element system composed by the pre-processor MTOOL 3D®, the simulator ANVEC 3D® and the post-processor POS 3D®. Salt rocks are modeled as visco-elastic materials. Creep is simulated by the Double Mechanism Creep Law, widely employed in Brazilian salt rock mechanics research. Cement and casing (steel) are modeled as elasto-plastic materials described by the Mohr-Coulomb and the von Mises criteria, respectively. Casing integrity is evaluated with respect to vertical and horizontal displacements. Vertical displacements imposed to the casing can provoke the failure of the column by plastic collapse or by buckling. Besides that, horizontal displacements of salt due to creep closure shall lead to the undesirable contact with the casing. Failure envelopes based on displacements are generated. The results of the simulations have shown that the casing failure by plastic collapse occurs preferentially in short columns, as well as buckling occurs in long and slender columns. The occurrence of both failure modes is more frequent in large-diameter cavities. Additionally, failure caused by contact between salt and casing occurs preferentially in short-diameter cavities.

Jan 1, 2015

By M. Sukhram

This paper presents a technique to determine the direction of velocity in liquid aluminum (Al) by measuring the temperature distribution within a cylindrical rod while it is inserted into the flow. The heating pattern of the rod is a result of the interaction of the thermal fields of the rod with the liquid Al. This interaction freezes aluminum onto the cylinder in a characteristic pattern and as a result the temperature response within the rod also follows a distinctive heating pattern. Monitoring the temperature within the cylindrical rod as a function of spatial and temporal coordinates, the direction of velocity can be inferred. Experimental research work involving liquid Al was conducted at the University of Toronto using the Revolving Liquid Metal Tank (RLMT). The RLMT is a resistance furnace which can hold up to 50kg of liquid Al. In this RLMT tangential velocities up to 0.35m1s can be generated. Experimental data for different superheat and magnitude of velocity demonstrate that the sensor records different heating patterns under different fluid flow conditions. Using the transient temperature information, the direction of velocity of the molten Al flow field can be detected.

Jan 1, 2011

The recovery of fine hydrophilic gangue in flotation is related to the recovery of water. Water is carried by bubbles as a film and wake into the froth. The amount carried (entrained) depends on bubble size, gas rate and the subject of this paper, frother type. To isolate the effect of frother type from that of bubble size and gas rate, gas holdup is used as the correlating variable. Measurements were made using a bubble column (no solids) operated continuously with the overflow recycled. Continuous operation meant some foam depth was required to ensure only entrained water is recovered. Using an automatically controlled 7cm foam depth the overflow rate of water is used as a measure of &apos;water carrying rate&apos;. It is shown that the common frothers can be grouped into four families based on the water carrying rate-gas holdup relationship.

Jan 1, 2006

By F. Hassani, E. Aflaki, M. Nokken, M. Kermani

"This paper presents and compares the physical properties of Gelfill and cemented hydraulic fill (CHF) obtained by lab experiments. Gelfill consists of tailings, different binders such as Portland cement, fly ash and blast furnace slag and an alkali activator such as sodium silicate. The CHF and Gelfill samples with various mixture designs were cast and cured for 28 days. The influence of sodium silicate concentration, pulp density and mixing time on the mechanical properties of samples were evaluated using the uniaxial strength test. Mercury intrusion porosimetery (MIP) was used to analyse the microstructure of CHF and Gelfill samples.This study concludes that: (i) The mechanical properties of CHF can be improved by the addition of appropriate amount of sodium silicate. (ii) The rate of strength acquisition in Gelfill samples is faster than CHF samples over 28 days of curing period. (iii) Mixing time strongly influences the mechanical properties of Gelfill samples but did not have any effect on CHF samples. (iv) The MIP analyses indicate that the pore size distribution and total porosity of Gelfill and CHF were different, and this did contribute to higher strength development of the Gelfill samples. INTRODUCTIONThe increased depths of mines in the Canadian Shield and the high stress associated with these depths favour a new method of mine backfill that can both meet strength requirements and help mine productivity. The mine backfill mainly consists of tailings, water and binder materials, the latter being most expensive. Gelfill is a new mine backfill material whose binder usually consists of alkali activators such as sodium silicate and other cementitious materials such as blast furnace slag and normal Portland cement. Although sodium silicate has been used in concrete manufacture, the use of this material in mine backfill is relatively new (Doucet, 2007, Razavi, 2007). Until very recently there have been only a few isolated publications, mostly out of McGill University, regarding mine Gelfill (Razavi, 2007 and Kermani et al., 2009 and 2010). These papers have investigated some of the basic mechanical properties and mechanical behaviour of Gelfill. However, the physical performance of Gelfill in various conditions has yet to be understood. Therefore, the main objective of this study was to investigate the influence of mixing time and sodium silicate concentration on the mechanical performance of Gelfill and CHF through a series of experiments. Moreover, the microstructure of the Gelfill and CHF samples was studied."

Jan 1, 2015