With the objective of developing a wet, low-intensity magnetic separator model, a number of plant data were analysed. Although the particle size versus mineral recovery relationship for a given stage of magnetic separation in a plant appeared to be reasonably stable, it was found that magnetic separation could not be classified as a deterministic process on this basis. Accurate modelling required size-by-size liberation information. Detailed analysis of available data led to the development of a model based on tdhe pseudo liberation concept, which could be used when such liberation is not available. The underlying idea is the fact that separation of gangue particles becomes more difficult and magnetite recovery increased, as the feed grade to a magnetic separator gets higher. Feed grade to a magnetic separator was considered as a compact parameter defining the liberation state of feed. A three-dimensional relationship between feed grade, particle size and mineral recovery was constructed using available data from a given plant, which allow calculation of a particle size mineral recovery relationship for a feed grade, hence liberation state. The model was incorporated into Usim Pac and used for plant simulations. Although the model was not fully validated by implementing simulation-based modifications in a plant, results of simulations appeared to be reasonable. (C) Institute of Materials, Minerals and Mining and Australasian Institute of Mining and Metallurgy. Published by Maney on behalf of the Institutes.
Thermal studies were performed on the segregation roasting of pure nickel sulphate. The reaction path was found to be the same as for oxide segregation, although some deviation was observed during product,formation, where a metallic nugget was obtained as the end product. The nickel content in the metallic nugget has been analysed at 74% (remainder sulphur). The optimum reaction temperature appears to be 1100degreesC; with 10% CaCl2 + 10% charcoal addition (by weight); after roasting for 30 min, it is possible to extract a product containing 97% metallic nickel with 3% sulphur. (C) 2004 Institute of Materials, Minerals & Mining and Australasian Institute of Mining and Metallurgy. Published by Maney on behalf of the Institutes. Manuscript received 20 August 2002; accepted in final form 24 June 2004.
Experimental and thermodynamic studies on the iron-sulphate-water system were carried out in order to appreciate the possible reaction mechanisms for the formation of ammonium jarosite. The mole ratio of ferric and sulphate species in the precipitates obtained during the kinetic studies was analysed. Thermogravimetric analyses (TGAs) and differential thermal analyses (DTAs) were conducted to examine the weight losses from ammonium jarosite during its thermal decomposition. Based on the results of both the thermodynamic studies and the experimental test work, a possible reaction mechanism for the formation of ammonium jarosite is proposed. This proposed reaction mechanism and the possible multiple pathways by which jarosite precipitation could occur are discussed. When ammonia solution is used as the ammonium ion source, the formation of ammonium jarosite is belieived to start with the species FeSO4+ and Fe(OH)(3). Solid basic ferric sulphates. [Fe(OH)(2)](2)SO4 and Fe(OH)SO4 may be precursors for the formation of ammonium jarosite Fe(OH)(3) can undergo reactions forming hydroxyl feric Fe(OH)(2)(+) ions and these are able to form basic ferric sulphates. There appears to be a gradual solid state conversion of the basic ferric sulphates/hydroxides into jarosite over time. (C) 2004 Institute of Materials, Minerals and Mining and Australasian Institute of Mining and Metallurgy.
A method is described for the removal of high arsenic levels from a gold concentrate involving a weakly reducing atmosphere in a rotary pipe furnace. The results showed that 95% of the arsenic in the gold mineral could be removed if roasted for 40 min at a temperature of 650-700degreesC with the resulting roasted materials containing less than 1% As. The concentrate also contained a high concentration of S (similar to20 wt%) and, under these operating conditions, over 25% of the S was also removed. (C) 2004 Institute of Materials, Minerals and Mining and Australasian Institute of Mining and Metallurgy.
A graphical method to assess the size reduction by grinding is presented and illustrated with data collected from laboratory grinding tests and industrial grinding circuits. The method lies between the graphical comparison of size distributions and the population balance model. The method is easy to apply and provides a quick way to compare grinding mill performances under various operating conditions and processed ores. The application of the method to industrial mills shows that a grinding mill operation could present an invariant behaviour that can be exploited for simulation purposes. (C) 2004 Institute of Materials, Minerals & Mining Published by Maney on behalf of the Institute in association with AusIMM. Manuscript received 28 November 2003; accepted in final form 5 June 2004.
Copper (last front ICC, Ghatsila (India) was characterised and water leached. Water leaching at room temperature for 1h dissolves as much as 50% copper present in the sample. The water leach residue, consisting mostly of oxides, was moistened with H2SO4 then baked (sulphatising roasted) in static conditions for 30 min at 973 K The baked mass was found to contain CuSO4, CuSO4.3H(2)O CuSO4.5H(2)O, Cu2O(SO4) and Fe2O3. Copper in the baked mass can be leached quantitatively in I h at room temperature by even 0(.)1% H2SO4 solution. Optimum conditions of baking and leaching for the highest recovery of copper have been determined The solutions obtained by sulphuric acid leaching of water leached and baked material contain about 2-3 g/l iron in both ferrous and ferric states. This iron can be largely precipitated out by adding lime to adjust the pH to 3(.)60 followed by passing air through the solution at 353 K for 4 h. The precipitate can easily be separated by filtration and the filtrate is of electrorefining electrolyte grade containing 44(.)05 g/l copper and 0(.)9 g/l iron. The iron content may further be reduced to 0(.)63 g/l by single stage extraction using 0(.)05 M D2EHPA solution.
Salt-type minerals of calcium (calcite) and barium (barite) are important raw materials for metallurgical. chemical and algricultural industries. Flotation is the primary process for recovering these minerals from ores. This study, reports the determination of the sit face roughness and wettability characteristics of these salt-type minerals ground in ball, rod and autogenous mills. Sin-face roughness values of these minerals have been expressed by the calculated surface roughness (R-BET)factors based on BET (Brunauer-Emmett-Teller) surface area measurements. The wettability characteristics (gamma(C)) of these minerals gound by three different mills were also determined by a microflotation technique using the EMDEE MicroFlot agitator Finally signi correlations were found between the surface roughness (R-BET) wettability; (gamma(C)) characteristics. The results revealed that the gamma(C) values increase with increasing R BET values. This means that if a mineral has lower surface roughness value, it will float better with lower gamma(c) value that indicates more hydrophobic character The empirical relationships could be stated as gamma(C) = a x R-BET + b type of equations: where a and b are constants. (C) 2004 Institute of Materials, Minerals and Mining and Australasian Institute of Mining and Metallurgy. Published by Maney on behalf of the Institutes.
Coal basins may have different coal formations. Therefore, different coal characteristics can be seen even in the same coal seam. The characteristics of the coal and the type of the mineral matter in coal are very important factors for the design of coal washing plants. The Soma coal region in Turkey has different coal formations which necessitate a careful blending of the coals prior to the washing plant. At present, Soma coals are being washed in a conventional heavy media plant. However, coals from different productions zones are fed to the plant without any blending. The clean products are also not blended to gain a consistent quality before sale. In this study, five different types of coal from the Soma region were washed together in the same washing plant. First, these coals were submitted to float-sink tests. Data derived from these tests were used for the evaluation of the plant trials. A novel method was used for the evaluation of the coal's washability and its probable effects on the performance of the coal washing equipment. As a result, experiments and evaluations show that the coal formations in the Soma region have significant differences in their characteristics which can impair the performance of the coal washing plant. Thus, washing these coals in the same plant under the same conditions needs blending or a special feeding programme. (C) 2004 Institute of Materials, Minerals & Mining Published by Maney on behalf of the Institute in association with AusIMM. Manuscript received 28 November 2003; accepted in final form 6 June 2004.
A beneficiation procedure for possible refractory uses of Bitlis Massif (Turkey) kyanite ore has been developed and involves grinding to d(80) = 0.106 mm, de-sliming, mica flotation, iron oxide flotation and kyanite flotation. A concentrate containing 56-5% Al2O3, 39-01% SiO2, 0-8% Fe2O3, 0-18% TiO2, 0.09% Na2O, 0-1% K2O and 0.02% MgO was obtained at a mass recovery of 51.14% +/- 0.25%. This conforms to both the chemical and size specifications for ceramic uses required to meet American and Chinese standards. (C) 2004 Institute of Materials, Minerals & Mining. Published by Maney on behalf of the Institute in association with AusIMM. Manuscript received 23 December 2003; accepted in final form 2 June 2004.
The incomplete Beta function was tested as a stochastic model to describe the compositional distribution of the particles obtained by grinding a real ore. A feldspatic rock, composed mainly of quartz, albite and microcline, was used. Polished sections were prepared from both unground samples and sized fractions of the ground material. The sections were measured by image analysis to determine the mineral quantities and the mineral distributions in the particles of the different size fractions. The composition distributions were used to test and validate the model Statistical analysis was done to evaluate the differences between the measured and the theoretical composition distributions, generated by the incomplete Beta function. For most fractions, the model did not pass the Kolmogorov-Smirnov test of adherence at a level of 0.05. At levels of 0.01 and 0.001 the adherence extended to a greater number of fractions. (C) 2004 Institute of Materials, Minerals & Mining Published by Maney on behalf of the Institute in association with AusIMM. Manuscript received 22 September 2003; accepted in final form 13 April 2004.
The flotation circuit of the Sarcheshmeh copper mine, which processes 41000 t day(-1) of ore tit 0(.)9% Cu, consists of rougher, cleaning and re-cleaning stages. To study the stage recovery of copper, one of the eight identical flotation banks was selected,for sampling It was observed that 17% of copper goes to the tailings and the lowest copper recovery was for particles coarser than 105 mum. The third 5-cell rougher unit of this bank not only did not recover an appreciable amount of copper but also decreased the concentrate graded The grade of the last five cells was lower than the feed grade suggesting that mixing of the concentrates of the three units was not beneficial. To alleviate the problem, the concentrate of the first 4-cell unit as directed to the cleaners' feed box (i.e. by-passing the re-grind mill). This increased the overall copper recovery by 1(.)7%. Also, rather than the current practice of adding the reagents only tit the feed tank (i.e. 100-0-0%), various combinations of stage addition were examined. It was found that a 75-25-0% addition regimen could increase the total copper recovery by 1(.)3%. At the re-cleaning stage in order to increase the final copper concentrate grade, a froth washing system was used. To ease the maintenance of the system, a shower type system was installed. The optimum wash water superficial flow-rate was found to he 0(.)06 cm s(-1). Continuous sampling of the re-cleaners with and without wash water over a period of 2 months showed an increase of 1(.)5% in the final copper concentrate grade tit constant overall copper recovery.
Cyanide spills and exposures in the mining industry have occurred in many countries and in recent times there has been extensive publicity, with some moves towards the banning of cyanide in mining. Cyanide spills or accidental release of cyanide-bearing tailings and solutions into the environment, mostly as a result of containment failure, are usually managed by detoxifying the residual cyanide using well known methods before allowing the escape of the sludge. Ghana, a small West African gold mining country, has not been spared these unfortunate incidences. In the last decade and a half, eight incidents of cyanide spills have occurred. The most recent, in October-2001, was at Gold Fields Ghana Ltd in Tarkwa. This paper chronicles the various cyanide spills in Ghana and makes all assessment of the recent spill in Tarkwa. It is noted that the cyanide detoxification process contributed in no small measure to the adverse impact on the environment.
The discrete element code PFC3D has been used to model lifter stresses within a large tumbling mill. The intensity of the induced stresses (shear and normal) is directly proportional to the intensity of lifting action and liner/lifter wear. Results show that, for the modelled case, the magnitude of the stresses decreases as the number of lifters increase. Hence, longer intervals between relining can be expected for a mill with a larger number of lifters. However, it appears that beyond a critical number, a further increase in the number of lifters will not result in significant further reduction of stress. Distribution of the impact energy is also affected by the number of lifters. With very few lifters, the dominant form of energy consumption will be low intensity abrasion events. With larger numbers of lifters, high intensity impacts will be more frequent. Net power draw will be at a minimum for very small numbers of active lifters. As the number of lifters increases, power draw increases and eventually reaches a relatively value. Further work will be required to investigate the effect of lifter shape on the induced stresses. It appears to be possible to determine the number and shape of lifters that will result in the optimal power draw, while simultaneously minimising liner/lifter wear.
Germanium is present as a minor element in most mined lead-zinc concentrates, and during subsequent processing part of it is carried in electrolyte to zinc electro-winning. Germanium is well known to reduce drastically the current efficiency of zinc deposition and to cause a multitude of pinholes in the cathode deposit, even at concentrations <0.1ppm, so it is important to understand how to minimise the amount of germanium proceeding to zinc calcines leaching. This work was aimed at determining the distribution of germanium during lead smelting among lead, slag and gas and generating thermodynamic data for use in process modelling. Lead and a small amount of GeO2 was equilibrated with CaO-SiO2-FeO-Al2O3 slag under oxygen partial pressures from 10(-10) to 10(-12.5) atm at temperatures from 1150 to 1250degreesC. It was found that most germanium is incorporated in the slag, some volatilises and little dissolves into the lead. Germanium is present in the slag predominately as GeO and its activity coefficient in this slag is calculated to be in the range 1.44-2.55. The implications of these findings on processing are also discussed.
Most processes of steel manufacture necessarily generate a substantial proportion of slag. Often much of this is landfilled, and at many plants it will be the main item (by tonnage) that is dumped. This rock-like material can be used as an aggregate, but the main obstacle to widespread use is concern over the volume stability. The main cause of volume instability is hydration of free lime, although hydration of free magnesia may also contribute in some cases. There are three key steps to effective use of this abundant secondary aggregated First, some pretreatment of the slag is usually necessary (such as weathering). Second, a test is necessary that reliably predicts the behaviour in use within a reasonable time. This is commonly an expansion test. Third, calibration of the test is necessary so that test results can provide a useful distinction between material which is suitable and that which is not; this is normally achieved by linking laboratory tests to road trials. Subject to these three steps, steelmaking slag can make a valuable contribution to the need for secondary aggregates.
The feasibility of recovering manganese by leaching low grade ore with aqueous sulphur dioxide has been investigated. Successful extraction ivas achieved at atmospheric pressure and room temperature. The concentration of manganese in the leach liquor increased with increase in pulp density. At higher pulp densities (e.g. 40 and 50%), the extraction of impurities Fe, Al, Si and Na into the leach liquor decreased, which was advantageous. A manganese rich leach liquor with 119 g L-1 Mn, 1(.)18 g L-1 Fe, 1(.)03 g L-1 Al, 1(.)65 g L-1 K, 0(.)33 g L-1 SiO2, 0(.)31 g L-1 Na and 0(.)23 g L-1 Co could be obtained from the low grade ore at 50% pulp density and 150 min leaching time.(C) 2003 IoM Communications Ltd. published by Maney for the Institute of Mateirals, Minerals and Mining in association with AusIMM.
Significant environmental benefits are secured in the gasification of coal, petroleum coke, biomass and similar carbonaceous materials, if the ash content is rendered non-leachable, by the production of a vitreous slag. By combining float glass and generic melt circulation technologies, a new, liquid-metal based process for gasification is proposed The first case considered focuses on low sulphur feeds for which the proposed basic flow-sheet involves a pressurised horizontal liquid-metal transport gasifier, to which solid charge materials are simply introduced onto the moving liquid metal surface via a barometric leg entrained in liquid metal. The solids are transported in a straight line through an air and steam top-blown zone to gasify, the solids and produce a liquid sheet of molten slag, analogous to float glass manufacture, except that the liquid-metal is force circulated through the gasifier and external heat and mass transfer functions, involving direct countercurrent contacting in irrigated packed or moving beds to remove solid particulates and gaseous pollutants at high temperatures In the context of advanced power generation, the fuel gas is cooled to around 240degreesC to permit straightforward filtration and then is re-heated again by direct liquid-metal contacting to at least 730degreesC for transmission to the combustors of a gas turbine, which yields enhanced combined cycle efficiencies. More complex flow-sheets are then explored in which metallic zinc is added non-intrusively in a separate gas circuit independent of the fuel gas circuit to effect desulphurisation and to yield a valuable by-product, zinc sulphide. It is shown that not only can a slag be produced in a single stage air-blown cycle without the need for air separation and oxygen production, but also toxic components in air, in general, can be removed; in the particular case of mercury, this can be recovered directly in elemental form.
Melt circulation and conductive heating arc introduced in relation to the then new imperial smelting furnace (ISF) for the production of zinc in a blast furnace. The focus then changes to projected development of a recently proposed continuous steelmaking process, in which pools of molten iron/steel are created inside residual solid ferrous material to form a cascade of interconnected melt circulation loops. These pools are maintained at the liquidus temperature by melt circulation with the thermal requirements being provided by conductive heating, during start-up, stand-by or planned shut-downs. During operation, the various 'swimming pool reactors' process well-mixed iron bearing materials (virgin ore or secondary material) with granular/pulverised coal and flux, deposited as charge on the moving surface of the iron-making loop. A solid raft of metallised material is first formed and then projected by drag forces onto the surface of the first of the steelmaking loops, in which melting and slag/metal separation occurs along with partial decarburisation and desulphurisation. The continuous overflow, by a hot metal siphon is treated in at least one further melt circulation loop to produce refined steel ready for continuous casting. Key features of hot metal siphons, emphasising the requirements for conductive heating, are discussed in detail.
The power required to operate large mills is typically 5-10 MW. Hence, optimisation of power consumption will have a significant impact on overall economic performance and environmental impact. Power draw modelling results using the discrete element code PFC3D have been compared with results derived from the widely used empirical Model of Morrell. This is achieved by calculating the power draw for a range of operating conditions for constant mill size and fill factor using two modelling approaches. fThe discrete element modelling results show that, apart from density, selection of the appropriate material damping ratio is critical for the accuracy of modelling of the mill power draw. The relative insensitivity of the power draw to the material stiffness allows selection of moderate stiffness values, which result in acceptable computation time. The results obtained confirm that modelling of the power draw for a vertical slice of the mill, of thickness 20% of the mill length, is a reliable substitute for modelling the full mill. The power draw predictions from PFC3D show good agreement with those obtained using the empirical model. Due to its inherent flexibility, power draw modelling using PFC3D appears to be a viable and attractive alternative to empirical models where necessary code and computer power are available.
The latest technology for processing Parkes silver crusts from lead refining uses vacuum indiction furnace retorting to remove zinc, followed by bottom blown oxygen cupellation for oxidising lead to produce silver dore suitable for electrorefining. A computational thermodynamics model of the cupellation process has been developed to predict the effect of operating variables on the final dore composition and silver recowry to dore, as an aid in optimising the silver production process. However, this model was limited by the absence of any thermodynamic data on AgO0.5 in solution in molten PbO. In the present work, the activity of AgO0.5 in molten PbO has been determined at 1000degreesC over the composition range of interest. It was found that the activity of AgO0.5 obeys Henry's law and that the limiting activity coefficient is 0.65. These data are used in the computational thermodynamics model to discuss some important relationships in bottom blown oxygen cupellation.
