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By A. Güney, G. Önal

"The tailings of the Bigadiç Colemanite Concentrator at the West of Turkey which contain 18 % B2O3 and 1.57 % SO3 are collected at the tailing dams near the Concentrator. In this research, it was aimed to obtain colemanite concentrates with a minimum content of 40 % B2O3 and a maximum 0.5 % SO3 from the tailings. At the experimental studies, decrepitation methods were systematically applied following attrition scrubbing and sizing. At the end of the tests, the best decrepitation result were described, at 500oC temperature for 15 min. duration, the final concentrate having 57.84 % B2O3, 0.44 % SO3 with 73 % recovery was obtained.Boron ores reserves of Turkey are about 2.5 billion tons and Turkey has the 63% share of the World’s boron reserves. The boron reserves are located at Eskisehir (Kirka), Kutahya (Emet), Balikesir (Bigadic) and Bursa (Kestelek) districts in Western Turkey. In 2002, the total amount of boron production of Turkey was about 1.4 million tons, which is 40 % of the World boron production.Bigadiç colemanite (Ca2B6O11. 5 H2O) ore with a content of 30-31 %B2O3 are concentrated by attrition tumbling and classification process. So the final concentrates with contents of 40 and 42 % B2O3 are obtained at the size fractions of –125+25mm and –25+ 3 mm respectively. The tailings materials with an 18 % B2O3 content are collected at the tailing dams near the Concentrator. In this research, it was aimed to obtain colemanite concentrates with a minimum content of 40 % B2O3 and a maximum 0.5 % SO3 from Bigadic tailings having 18 % B2O3 and 1.57 % SO3.In experimental tests, attrition scrubbing, wet sieving, gravity separation, electrostatic separation, flotation and decrepitation were adopted to the colemanite tailings, and optimum beneficiation conditions were obtained."

Jan 1, 2003

By Ashton M. Swartbooi, Robert van Hille, Alison E. Lewis

"Nickel production by precipitation using hydrogen reduction generates an ammonium sulphate solution containing approximately 2g/L of nickel, and 1g/L of cobalt as an impurity of the process. Sulphide precipitation is an effective method for heavy metal removal, especially when compared to hydroxide processes: metals can be removed from solution at relatively low pH levels, small quantities of sludge are formed, and due to the high reactivity of sulphides the reactor retention time is reduced. However, the low solubility of sulphide species can lead to spontaneous nucleation of fines (due to higher supersaturation levels which are difficult to separate from solution, leading to reduced process efficiency. Using a fluidised bed reactor, fines formation can be minimised as the pellets provide a large surface area and promote heterogeneous nucleation.A synthetic nickel-cobalt ammonium sulphate solution was reacted with two forms of sulphide reagent: (a) a liquid sulphide solution and (b) gaseous hydrogen sulphide. A range of factors (including pH, reagent concentrations, metal to sulphide molar ratio, flow rates and recirculation rates) influencing the efficiency of the precipitation reaction have been studied. The efficiency of the precipitation reaction was measured using both metal recovery and fines formation as indicators.Figure 1 shows the effect of decreasing the metal:sulphide ratio (at constant SS = 2.725e+17), on the metal removal for liquid sulphide. By decreasing the metal:sulphide ratio, up to 99% removal of both nickel and cobalt could be achieved.These results also indicate that the precipitates that formed were polysulphides, since no sulphide was present in solution after the experiment. This would also indicate that a lower metal:sulphide ratio would be more favourable. However, decreasing this ratio (at constant metal concentrations) would lead to an increase in the fines concentration, because of increased supersaturation levels induced by the higher sulphide concentrations. To decrease the supersaturation, dilute solutions of either reagent would be employed."

Jan 1, 2003

By Remmerswaal, U. Boin, M. B. G. Castro

"To design sustainable products that can be recycled with high resource efficiency, low environmental impact and economic feasibility, the choice of materials and materials combinations, made in the design stage, is fundamental. The products we design, manufacture and use now will be the resources of the future. Designers are able to select materials for a product according to the many functional requirements they must fulfil. But they have a limited knowledge of the recycling processes and its requirements regarding quality of the secondary material streams. Resource Engineers have to manage the processing of these streams, that are becoming more complex with the increasing variety of materials and decreasing size of components. After mechanical separation of the materials streams resulting from shredding of EOL (endof- life) products, there is always a degree of contamination, from two possible causes:Incomplete liberation: the bonds between different materials combined in a component are not completely destroyed by shredding, resulting in some particles containing two or more materials attached;Separation efficiency: the mechanical separation processes are not able to separate perfectly the many material particles; there are always a number of foreign particles in the recovered streams.Sometimes, those contaminants cannot be removed by the industrial metallurgic recycling processes, and a more serious problem occurs when the contamination levels compromise the recycling of the material stream by lowering its quality (i.e. the properties of the recycled material will be lower due to the contamination). As a consequence, the material must be diluted in virgin raw material and/or used in a less demanding application (downgrading), causing resource depletion and increased energy use, waste production and recycling costs. In order to make the best choice of materials combinations for a product, a systematic approach is necessary, that can close the gap between the knowledge areas of Product Design and Recycling Technology."

Jan 1, 2003

By Steven Wright, Sharif Jahanshahi, Shouyi Sun

"While there is a large industry based on recovering platinum group metals (PGM) from complex sulphide ores containing Cu and Ni, there is little information on the thermodynamics of the matte phase formed on smelting the ore and in particular the behaviour of Pt in the matte.The equilibrium between metallic platinum and industrial mattes was studied in a series of experiments where a small quantity of matte (5g) and platinum foil of 150 µm thickness (0.5 g) were equilibrated in magnesia crucibles at 1400°C under conditions where the oxygen and sulphur partial pressures were controlled.The chemistries of the phases were determined from EDS analysis using a scanning electron microscope. Line scans were initially used to confirm that the foils had attained a uniform composition across their thickness. The composition of the matte was determined from analysis of an area of matte including the small metallic droplets and dendrites.The distribution of Pt in Cu-Fe-Ni mattes (Pt: 1.5 – 3%) and solid Pt alloys (Pt: 54 – 72%) was studied. For a matte of base composition 45%Fe, 11%Ni, 8% Cu, at a base temperature 1400°C, the following behaviour was observed for the Pt distribution ratio:Under the range of sulphur and oxygen pressure of this study (log pO2 from -15 to –12.2, log pS2 from – 4.6 to –2.2), the distribution ratio of Pt between matte and alloy was independent of the sulphur and oxygen pressure, having a value of 0.04.The Pt distribution ratio increased from 0.02 to 0.08 with temperature from 1340 to 1450 oC.For mattes where the proportions of Cu and Ni were kept constant, a slight decreasing trend was shown for the distribution ratio of Pt versus the iron content (37 to 53%) in the matte but was not confirmed because of experimental uncertainty."

Jan 1, 2003

By P. Saha, D. M. Chakrabarti, National Metallurgical Laboratory, P. Bhattacharyya, B. R. Sarkhel

The estimated reserve of the phosphatic soils at the Sukulu hills region of Uganda upto a depth of 60 meters is of the order of 200 million tonnes and analyses 10-12% P2O5. A 5-tonne sample of the phosphatic soil was received at the National Metallurgical Laboratory (NML), India for beneficiation studies. The ore occurred as a naturally degraded soil deposit and the slime contained earthy lateritic minerals which made it extremely sticky. Microscopic examination revealed the presence of limonite, magnetite (martitised), hematite and goethite as major minerals along with apatite and quartz. Presence of clay was indicated by earthy smell on moistening the sample. It was observed that getting rid of the slime was as important for adequate beneficiation of the ore, as it was difficult. Earlier studies on samples of the same ore at the NML and abroad established that it was necessary to pass the ore through a series of three stages of desliming, each consisting of a step of rigorous attrition scrubbing followed by hydrocycloning, for effective elimination of the sticky slimy coating which stubbornly adhered to mineral grains. Studies were carried out at the NML in two phases to develop a flowsheet for beneficiation of the phosphatic soil. In the first phase, reproducibility of the results obtained by employing a flowsheet developed abroad was confirmed along with suggesting some modifications of the flowsheet. Then 200 kg. of concentrate was produced following the recommended flowsheet for evaluation of its suitability for production of fertilisers. To remove the limonitic slimy coating, which stubbornly adheres to the other mineral grains, it was necessary to subject the ROM ore, after screening out the +840 micron size fraction which contained only 1.75% P2O5 , to three stages of hydrocycloning each preceded by attrition scrubbing. Then only conventional beneficiation techniques could be effectively employed to upgrade the ore.

Jan 1, 2003

By M. Ghiani, R. Peretti, S. di Giovanni, G. Zonnedda, A. Serci

"An interesting occurrence of gold bearing ores of economic interest has recently been discovered in the Serrenti- Furtei district, central-southern Sardinia (Italy). The upper part of the orebodies are leaner and fairly porous, and are undergoing intense oxidation processes. In the lower parts sulphide ores predominate, chiefly enargite-luzonite and pyrite and these have a more compact structure. The passage from the oxidized to the sulphide zones, is rather irregular and difficult to delineate. The gold, generally occurring in its native form, is found in small grains. Mineral reserves in the area concerned were originally estimated to total 2.77 Mt: 0.63 Mt of oxide minerals with average gold content of 2.75 g/t and 2.14 Mt of sulphide ores with average gold grade of 3.18 g/t.Recovery of the valuable metals is a complex problem.The effectiveness of a variety of gold extraction techniques has been evaluated conducting tests on batches of ore representative of the different orebody depths obtained from core samples.The techniques investigated were: heap and vat leaching; flotation and flotation followed by leaching of the flotation tailings. The findings indicate that:• the gold contained in the ores in the upper oxidized zones of the deposit can be efficiently recovered by direct cyanidation via heap or vat leaching;• the sulphide ores in the deeper portions, that are not amenable to direct cyanidation but contain enargite concentrations of potential economic interest can be processed by means of flotation to produce a gold/copper bulk concentrate;• beneficiation of the difficult-to-treat minerals contained in the transition zone requires some knowledge of their intrinsic properties. A distinction should be made between the ores containing minor amounts of enargite which appear to be amenable to vat leaching and those with relatively higher enargite content which are processed using bulk flotation of the gold and copper minerals followed by vat leaching of the resultant tailings."

Jan 1, 2003

By Olgac Kangal, Ali Guney, Gulay Bulut, Guven Onal, Ayhan A. Sirkeci

"This paper presents the results of pilot scale column flotation applications in the two different copper concentrators in Turkey. The Kure Copper Plant is located in the Black Sea Region of Turkey. There are four types of copper ore formations in the Küre region. The so called disseminated ore contains 60-70% siliceous minerals of which Cu content is around 0.5%. The remaining three ore types in the region are generally named “massive ores”. The massive ores may assay up to 15% Cu and the major gangue mineral is pyrite associated with relatively little amount of silica. The main copper mineral in all formations is chalcopyrite that is generally finely disseminated in pyrite therefore satisfactory liberation requires ultrafine grinding especially in the case of the massive ores.A blend of these ores, containing 1.7% Cu on average, is fed to the Küre Copper Concentrator whose capacity is 1.2 million tons of ore annually. Unfortunately, today the plant operates below its designed capacity treating around 550,000 tons of ore per year. The design of Küre Copper Plant was undertaken by the Outokumpu Company of Finland. The operation of the plant is almost automatically controlled and the feed, the concentrate and the tailings are assayed for copper by on-line XRF analysers. The particle size analysers are also connected to the grinding circuit. The feed material is ground almost below 50 µm and the flotation is conducted around pH 11.2 using Hostaflot X-231 (a trade mark of Hoechst) as the collector. Apart from the rougher cells, there are four cleaning stages after which the final concentrate is produced. A very fine grinding requirement for an adequate level of liberation makes successful flotation difficult and consequently the final concentrate can hardly assay above 16% Cu leading to copper losses around 20% in the tailings that contains 0.3% copper."

Jan 1, 2003

By O. N. Mohanty, S. Mohan Rao, N. D. Rao, T. Mukherjee, P. V. T. Rao

"Chromite ore is an indispensable raw material in the manufacture of ferro-chrome and charge-chrome, which are used for the production of stainless and other alloy steels. The demand for chromite ore is ever increasing because of increasing consumption of these steels. Keeping this in view, Tata Steel is keen to recover values from fines and wastes, and also to produce value added products from their captive chromite mines and its processing plants.Tata Steel has its captive chromite resources at Sukinda, Orissa. They contain two ore types, i.e. (i) Ferruginous and (ii) Siliceous. The generation of off-grade ores is an increasing trend and simultaneously customer quality requirements are becoming ever more stringent. The ferruginous type ore also contains a considerable quantity of in-situ fines (~20% particles below 0.5 mm) that compound the beneficiation problem.In many places, in India and abroad, mined non-marketable off-grade and low grade chromite ores are upgraded by using either simple classification techniques like screw classification, hydrosizing, hydrocycloning or conventional gravity concentration processes such as Heavy media separation, Jigging, Spiraling, Tabling etc.At Tata Steel, the ferruginous type ores are treated in Chrome Ore Beneficiation (COB) Plant that comprises: a crushing and grinding circuit; a classification circuit using hydrocyclones; and a beneficiation circuit using 4 stage spirals in the coarse circuit (-0.5+0.1 mm) and 3 stage spirals & tables in the fine circuit (-0.1+0.038 mm). In this process, 3 types of tailings are being generated and together analyses about 17 % Cr2O3, which is high. From the present operation the chromite recovery is around 70% and could not produce coarse concentrates (+0.5 mm size). Whereas the siliceous type ore is treated in the Lumpy Ore Processing (LOP) plant, consisting of crushing-washing & classification-hand sorting operations, to produce marketable grade lumpy ore and –10 mm de-slimed fines. These fines are intermittently treated in the COB plant. The higher silica content (greater than 1%) and low Cr/Fe ratio (less than 2.8) are the problems with these concentrates."

Jan 1, 2003

By Xin Tan, Changgen Li

"The floatabiliby of chelate CF-collector for zinc and lead oxide minerals was described. It found that CF-collector possesses a strong collection power for zinc and lead oxide minerals as well as excellent selectivity for the gangue minerals. A new technology of zinc and lead oxide flotation with CF reagent as a collector, lead nitrite as a activator and water glass as modifier was presented. The mechanism of interaction between reagents and minerals was discussed by means of adsorption, zeta potential, nephelometric titration, XPS and infrared spectra was investigated.CF-collector, newly developed by Beijing General Research Institute of Mining and Metallurgy (BGRIMM), is a kind of chelate collector for the flotation of oxide minerals. CF-collector mainly consists of N-nitroso-N-phenylhydroxylamine ammonium salt. The CF-collector was used for the selective flotation of scheelite in the Chinese plants.In this paper, the effects of the dosages of activator Pb(NO3)2 and depressants (sodium hexametaphosphate, polyphosphate, water glass and salified water glass) on the floatabilities of several minerals (cerussite, smithsonite, calcite, dolomite, limonite and quartz) as well as their mixtures were investigated in the presence of the CF-collector. The results show that the CF-collector is the excellent collector for selective flotation of cerussite and smithsonite. Pb(NO3)2 can selectively activate the oxides of lead and zinc resulting in a decrease of the CF-collector dosage. The sodium hexametaphosphate plus the salified water glass are the effective depressants for gangue minerals. Based on the above facts, the “CF process”, namely, used CF reagent as a collector, Pb (NO3) 2 as a activator, and sodium hexametaphosphate plus salified water glass as depressant, has been used for the flotation separation of the lead oxide mixtures or zinc oxide mixtures without disliking and heating. For the feed lead assaying 19.07%, the lead concentrate assaying 49.81% Pb at the recovery 80.69% was obtained after one step flotation. For the feed zinc assaying 12.85%, a zinc concentrate of assaying 29.56% Zn at recovery 83.87% was obtained."

Jan 1, 2003

By Jean-Paul Franzidis, Robert G. Coleman, Emmy V. Manlapig

"The Floatability Characterisation Test Rig (FCTR) is a self-contained, highly instrumented mobile pilot flotation plant developed by the Julius Kruttschnitt Mineral Research Centre, the Mineral Processing Research Unit of the University of Cape Town and EIMCO, as part of the Australian Minerals Industries Research Association (AMIRA) P9 Project. Apart from being a fully operational pilot plant, the FCTR was designed to determine flotation model parameters and to develop and validate flotation plant modelling, scale-up and simulation methodologies.The FCTR was installed at WMC’s Kambalda Nickel Operation where circuit modelling, simulation and scale-up test work was performed using the AMIRA P9 flotation model as the basis of the analysis. The model parameters were obtained through metallurgical sampling, batch flotation tests and gas dispersion characterisation around the FCTR. Model parameters were determined for four FCTR circuits, each of increasing complexity. The circuits ranged from a simple rougher-cleaner configuration with no recycle streams, to a more complex rougher-cleaner-recleaner configuration with two recycle streams.Using the P9 modelling methodology, the system was analysed using four mineral species: nickel sulphide, iron sulphide, copper sulphide and non-sulphide gangue. The floatability of each mineral was analysed as two floating species (fast and slow) and a non-floating species. Mass balanced data from all four FCTR circuits were regressed simultaneously together with data from 47 batch flotation tests. Nonlinear regression was used to determine the solution. The model fit was found to be accurate with most data points predicted within experimental standard deviations. A Monte Carlo sensitivity analysis was also performed to establish the stability of the regressed parameters.Using the P9 simulation methodology, simulations of five FCTR circuits (different from the previous four) were performed. The FCTR was also operated in these five circuit configurations to obtain experimental data to validate the simulated predictions. The simulations were found to predict the recovery and grade of nickel sulphide and copper sulphide within experimental standard deviations. The prediction of recovery and grade of iron sulphide and non-sulphide gangue was less accurate. Overall the FCTR simulations predicted the changes in circuit configuration well."

Jan 1, 2003

By Markus A. Reuter, Jacques J. Eksteen

"The modelling of bath type furnaces tends to be either fundamental in nature or total empirical ""black box models"". While models of the first kind are robust, they tend to lack accuracy, in cases where they are to be used for process control, and may take an overly long in development time. On the other hand, black box models tend to lack robustness of prediction for all circumstances and tend to be limited to interpolate based on historical data. While this may very often provide short-term accuracy, the models may have problems with stability in the long term. Moreover, the model structure may not reflect any of the fundamental properties of the system and remains opaque to fundamental analysis. The aim of the model development is in obtaining a model that have robustness, high accuracy, a short development time and which is adaptive to a changing environment. The models are specifically developed for on-line decision support or control of furnaces. The modelling approach is evaluated for a chromite bath smelting furnace. The role of systems analysis, mass balance smoothing, thermodynamic analysis are integrated into one generic algorithm for bath type furnaces. Not only are fundamental process relationships such the mass and energy balances satisfied through a scientific method of data reconciliation; the proposed modelling approach allows adaptive prediction, the first step towards predictive control or metallurgical decision support.It is shown that neither an ARMAX model only, nor a purely equilibrium model only, could predict the dynamics of the process very well. However, a combination of the two approaches more than doubled the accuracy as measured by the R2. It was noted for carbon dissolved in the alloy, that the improvement in R2 in using a hybrid modelling approach was more than the sum of the R2 of the individual traditional approaches. Similar improvements in modelling performance were noted for all other species.It shows further that process dynamics may totally overshadow the effect of thermochemical equilibrium (even when the thermochemical equilibrium prediction includes the recycled alloy and slag heels as additional feeds). The exclusion of process dynamics therefore leads to erroneous results."

Jan 1, 2003

By A. N. Zagorodnyaya, T. N. Bukurov, Z. S. Abisheva, A. S. Sharipova

"This work presents the results of the development of a flowchart for the processing of copper production dusts resulting in the production of of lead sulphate, tribasic lead sulfate (TBLS), ammonium perrhenate, zinc, copper-cadmium cake.Dust having the following composition (wt. %): Pb 39.3, Zn 6.5, Cu 5.5, Stotal 10.8, As 1.05, SiO2 0.6, Cd 0.8, Bi 0.04, Ag 0.02, Re 0.03 was used in the work. Chemical and X-ray analyses indicated that lead, zinc and copper are generally present in dusts in the form of sulfates, various amounts of oxides, sulfides, silicates; sulfur (according to chemical analysis data) is present in the form of sulfates (~ 90%) and in the form of sulfides (~ 10%).Studies have been made on two-stage stripping of copper production dusts containing lead, zinc, copper, cadmium, rhenium in the form of various compounds, at first with soda solution, then with nitric acid solution. It has been indicated that the combination of the operations allows the recovery of 98% rhenium, 82% lead, 92% copper, and 96% zinc and cadmium each in the solution. Furthermore, at the stage of dust leaching with soda, 80% of rhenium and 0.01% of nonferrous metals each are recovered in the solution. Remaining quantities of all metals are recovered in nitric acid solution up to maximum values. Such behaviour of metals is caused by metal phase content in dusts and properties being formed during leaching of compounds. At the stage of dust leaching with soda solutions, carbonates with various compositions are formed due to the exchange reaction between nonferrous metal sulfates and sodium carbonate. The conversion of PbSO4 into PbCO3 is apparently occurs in the solid phase and the conversion of readily watersoluble zinc, copper, and cadmium sulfates into that of carbonate salts occurs in the liquid phase. Rhenium passes into solution due to the presence of rhenium greater quantity in dust in the form of readily soluble perrhenate. Some portion of the perrhenate is, probable, blocked by lead sulfate and liberated in the conversion of PbSO4 into PbCO3."

Jan 1, 2003

By Jaime Sepúlveda

"A broad diversity of critical operating conditions may influence, in very convoluted ways, the actual performance of any given mineral processing system, to the extent that it becomes nearly impossible to empirically determine the independent, decoupled impact of each condition. An effective approach to better understand the behaviour of such complex systems is via the extensive use of sufficiently detailed and reliable mathematical models, allowing the process analyst to explore and learn about the independent effect of all relevant operational conditions. Certainly, these tasks are greatly facilitated if the appropriate software tools - such as the Moly-Cop Tools software package - are readily available. Moly-Cop Tools 3.0 is a set of 63 easy-to-use EXCEL spreadsheets designed to help process engineers characterize and evaluate the operating efficiency of any given grinding circuit, following standardized methodologies and widely accepted evaluation criteria. Moly-Cop Tools 3.0 covers conventional and semiautogenous mills, crushers, HPGR’s, rod mills and vertimills, with the corresponding mass balance and parameter estimation routines; plus complementary spreadsheets regarding Bond’s Law calculations, grinding media selection and performance evaluations and other useful utility functions. As a relevant example of application, specifically focusing on semiautogenous grinding operations, the current publication describes how the models can help us learn and rationalise critical empirical observations such as the relationship between feed particle size and optimal ball filling, the role of the rocks as autogenous grinding media as well as the impact of feed size and discharge grate opening on the mill grinding capacity. Over the last 20 years, Moly-Cop Tools has gained wide acceptance around the mineral processing community in recognition to the quality of the mathematical models contained, their simplicity of application and ease of sharing of the calculations amongst all interested parties, since it is available free-of-charge with no restrictions to make as many copies as needed."

Jan 1, 2014

By S. K. Kawatra, T. C. Eisele

"Coal-fired electrical power generation plants produce 56% of all the electricity in the U. S., burning approximately 860 million tons of coal annually. This results in production of 58 million tons of fly ash; 19 million tons of bottom ash and boiler slag, and 23 million tons of desulfurization sludges. These are a considerable disposal problem, and it would therefore be very valuable to be able to sell these residues and byproducts as useful materials, rather than paying to dispose of them in landfills. However, utilization rates are currently very low, below 32% even for the most thoroughly commercialized combustion byproduct (fly-ash). This poor utilization rate is a result of the high impurity content of combustion byproducts, which make them unsuitable for existing markets. The experiments described in this paper were therefore carried out to study the application of mineral processing techniques to upgrading combustion byproducts.Experiments were conducted to upgrade and purify low-grade combustion products to produce highgrade, readily saleable materials. A process was developed for removing unburned carbon from coal fly-ash by froth flotation, using reagents originally developed for flotation of oxidized coals. This technique could reduce the carbon content of fly-ashes containing up to 11.1% carbon to less than 2%. The froth product was sufficiently carbon-rich that the recovered carbon has value as fuel, or as low-grade activated charcoal.Fly ash was also examined as a binder for producing iron ore pellets, primarily as a substitute for bentonite. It was found that Class F fly ash in combination with an activator (calcium hydroxide) was an effective pellet binder at dosages comparable to the currently-used dosages of bentonite clay. While the flyash worked very well alone, tests in combination with bentonite showed that fly-ash/bentonite mixtures had significantly worse binding performance than either pure bentonite or pure fly-ash-based binder. This was determined to be due to the fact that bentonite and fly ash have profoundly different binding mechanisms, with fly-ash undergoing a pozzolanic chemical reaction while bentonite binds through a physical attachment process."

Jan 1, 2003

By Lokesh K. Thakur, Peter Radziszewski

"For the cases tested (11 mills with 15 operating conditions), the Bond abrasion test steel media wear rate prediction gives an overall error of 73% and a standard deviation of 192.5%. However, total steel media wear in a given mill (ball or SAG) grinding process is a product of three recognised wear mechanisms -- impact, abrasion, and corrosion. A total media wear model can be defined on the assumption that the effect of each wear mechanism can be independently determined which allows for a total media wear model to be defined as the summation of the wear result of each mechanism (eqn. 1). It is assumed that impact and abrasive media wear is a function of the energy dissipated in impact and abrasion as estimated by a DEM charge motion simulation. Corrosive media wear is assumed to be a function of the media surface area present in a mill.The 3 initial tests developed (steel abrasion wheel, corrosion batch mill, impact chamber) coupled with the impact and abrasion energy spectra from a DEM charge motion simulation produced total steel wear rate results for the same cases as the Bond abrasion test giving an overall error of 0% and a standard deviation of 66.1%. However, if one were to consider the general observation that wear is proportional to energy and maintain the assumption that corrosion is proportional to media surface area then one would realize that the total media wear eqn. (1) can be reduced to eqn. (2). As this is a linear equation and one has, for the same mill, media wear and operating data for three different operating cases, one obtains three equations with three unknowns. This defines a simple matrix manipulation to determine the constants k1, k2 and k3. From the previous data set, we had 6 identical mills grinding the same ore but having different worn liner profiles and charge volume which provided the opportunity to average out the ki values over 20 combinations of 3 cases. With the determined ki values, it was possible to determine a second estimate of the contributions of impact, abrasion and corrosion to total media wear giving an overall error of 0% and a standard deviation of 56%."

Jan 1, 2003

By Herman D. Mendoza

"The removal of finely dispersed mercury from the effluent of small-scale gold mining operations and the eventual recycling of the recovered mercury was investigated. The conventional method of amalgamation, or “ball milling”, involves the mixing of a substantial amount of mercury with ore to recover free gold during the grinding stage. During grinding, a portion of the mercury is transformed into minute droplets, which readily float and disperse in the effluent. A kilo of gold ore containing 20-30 g Au/t, when ground with 3 g of mercury, yields 0.005 to 0.050 mg/l of mercury in the effluent, exceeding the concentration of 0.002 mg/l allowed by law for surface water.A rapid recovery/ recycling process was introduced using the surface-chemical properties of finely dispersed mercury in aqueous suspension. It was observed that fine droplets of mercury in an aqueous environment readily attached onto a collector having positive surface potentials for pH 3 to 10. An example of a suitable collector is alumina (Al2O3), which has an iso-electric point of pH 9.8. When an effluent, coming from a settling procedure after an amalgamation-grinding stage, was passed through a column-bed packed with alumina fiber, removal and subsequent collection of finely dispersed mercury occurred under acidic conditions, pH 3-6. Recovery was around 99.5%, yielding an acceptable quality of water safe for disposal into the aquatic environment. The collected droplets of mercury, which are left in the column-bed of alumina fiber, were subsequently treated with a cleansing solution at alkaline conditions pH 10 -12. At these conditions, both mercury and alumina fiber carry negative surface potentials, which results in the eventual detachment of attached mercury from the alumina collector. The effective recovery of attached mercury from the alumina collector by the detachment procedure is only 30-35%. The collected mercury from the cleansing stage can be collected, cleaned and can be used for subsequent amalgamation processes. The attachment and detachment behaviour of finely dispersed mercury onto and from the microscopic collector alumina is shown in Figure 1. The result of attachment and detachment were analyzed using existing interaction energy theories like the Darguin-Landau- Verwey- Overbeek (DLVO) Theory."

Jan 1, 2003

By Xiong Dahe

"SLon vertical ring and pulsating high gradient magnetic separator utilizes the combined force field of magnetism, pulsating fluid and gravity to beneficiate weakly magnetic minerals with the advantages of large beneficial ratio, high recovery, matrix not easy to be blocked and excellent performance.We have developed SLon-750; SLon-1000; SLon-1250; SLon-1500; SLon-1750 and SLon-2000 VP HGMS. Now more than 200 of them have been successfully applied in China.Application in Qi Dashan Mineral processing Plant of Anshan Iron and Steel Company: The iron ore consists of magnetite and hematite. In 2001, six SLon-1750 magnetic separators are applied to control the tails grade of fine hematite. The results of flowsheet is: feed grade 30.15%Fe, concentrate grade 67.00%Fe, iron recovery 75.86%. The results set up the historic high record of this plant and the processing coast is greatly reduced.Application in Gushan Iron Mine of Ma Anshan Iron and steel Company: There are 6 SLon-1750 and 4 SLon-1500 magnetic separators applied in this mine to treat refractory hematite ore. The feeding grade is about 45%Fe. The final iron concentrate grade is about 60%Fe and the iron recovery is about 76%. The results are much better than the previous gravity flowsheet.Application in Meishan Iron Mine of Shanghai Baoshan Iron and Steel Company: The mined ore consists of magnetite, hematite and siderite. Because of a relatively high content of phosphorus, the iron concentrate could not meet the market requirements. Now 18 SLon-1500 magnetic separators are applied to reduce phosphorus, which is reduced from about 0.4%P to 0.246%P. The quality of iron concentrate is greatly upgraded.Application in Panzhihua Ilmenite Processing Plant: In the past, the -0.045mm ilmenite fraction was discarded in to the tails because of lack of high efficient equipment. In 2001, the plant built up a magneticflotation flowsheet to recover the fine ilmenite from the -0.045mm fraction with 6 SLon-1500 magnetic separators. The total results are: feed grade 11%TiO2, concentrate grade=47% TiO2, recovery 44%. Each year the flowsheet recovers 60,000t fine ilmenite concentrate."

Jan 1, 2003

By Buks Kruger

"After the milling and flotation of the platinum bearing ore, the flotation concentrate is electrically smelted in 6-in-line electric arc furnaces and the resulting furnace matte treated in Peirce Smith converters to remove the iron sulphide. The iron free (less than 1% Fe) converter matte is then granulated and sent to the Base Metals- and Precious Metals Refineries for further processing.During the smelting and converting processes two distinct gas streams are produced.The Peirce Smith converter off gas contains 4% - 9% SO2 which is treated in a conventional Sulphuric Acid Plant to produce sulphuric acid (H2SO4).The electric furnace off gas contains 0.9% SO2 and was (until September 2002) vented through a 91 meter high stack to atmosphere.Off gasses which contain less than 3% SO2 is difficult to handle and quite a few processes had been developed over the years to cope with this gas. A broad spectrum of process suppliers and technology approaches is available and the different technologies can be divided into a few main approaches i.e.Scrubbing of the off gas with an alkaline solution to produce an environmentally friendly and stable waste product which can be discharged onto a tailings dam or permanent landfill site.Physical absorption of the SO2 into a chemical solution to produce high quality chemically pure process chemicals.Catalytic reaction of SO2 with O2 to produce SO3 which is absorbed in water to produce sulphuric acid of less than 90% strength.In October 2002 Impala Platinum commissioned a Sulfacid Plant at its Smelter in Rustenburg which recovers the SO2 from the electric furnace off gasses into a weak sulphuric acid solution.The Sulfacid Plant technology is widely used in the paint pigment industry. This is, however, the first application of the technology on electric furnace off gas.The furnace off gasses contains flue dust which is removed in electrostatic precipitators. The gas exiting the precipitators normally still contains between 40 mg/Nm3 and 60 mg/Nm3 dust and passes through a quench-scrubber before entering the reactors. The quench-scrubber cools the electric furnace offgas down from ± 200 °C to 60 °C and act as a final dust removal stage."

Jan 1, 2003

By Glasswell K. Nkonde, Fredrick Chileshe, Steven Simukanga

Improving food security by increasing crop output is a primary objective for most countries in the sub- Saharan region. Most soils in the region are deficient in phosphorus an important nutrient after nitrogen in plant growth. Phosphates are therefore used together with nitrogen and potassium based fertilizers in improving crop production. Zambia like many other countries in the sub-Saharan region imports phosphate fertilizers for use by its farmers. However, indigenous phosphate deposits of igneous origin exist in Zambia of which the main ones are Chilembwe in Petauke district, Mumbwa sugarloaf in Mumbwa district, and Nkombwa Hill in Isoka district. If commercially exploited, these deposits could become a ready source of inexpensive phosphate for the local farming communities in the country. Previous research has shown that partially acidulated phosphate rock (PAPR) produced using Chilembwe phosphate rock is just as agronomically effective as the imported superphosphate fertilizers on food crops in most of the Zambian soils. PAPR is produced by reacting finely ground phosphate rock with only a fraction (usually 50%) of the stoichiometric quantity of the sulphuric acid required to fully convert the apatite in the rock to monocalcium phosphate thus making the process cheaper in terms of acid consumption than the poduction of single superphosphate where the apatite in the rock has to be fully converted to monocalcium phophate. The factors that affect the percentage of available P2O5 in the PAPR produced include particle size of phosphate rock, degree of acidulation, concentration of acid, process temperature profile, residence time and drying temperature. These factors were studied in a 25kg/hr PAPR pilot-plant in the School of Mines at the University of Zambia using Chilembwe phosphate rock. The PAPR pilot-plant consisted of a rotary drum granulator as the main reactor, a rotary drum dryer and a triple deck vibrating screen. Ground phosphate rock was reacted with sulphuric acid and water in the granulator to form a granular product which entered the drier. The dry product was screened into three size fractions. The 1 – 4 mm size fraction was the final product whilst the –1mm size fraction was recycled. The +4mm size fraction was sent to a crusher while the crushed PAPR was sent back to the screens.

Jan 1, 2003

By Renato Oliveira, Walter Valery, Alex Jankovic, Birol Sönmez

"The ball mill is the most common ore grinding technology today, and probably more than 50% of the total world energy consumption for ore grinding is consumed in ball mills. On the other hand, high pressure grinding roll (HPGR) technology is relatively new to the mining industry, with significant application only since the beginning of the 21st century. Both ball mills and HPGRs are typically used in closed circuit with classifiers. There is a general consensus within the industry that classification efficiency and circulating load both have a major effect on the efficiency of closed circuit ball mills. Metso Process Technology & Innovation (PTI) conducted a laboratory test program to understand the relationship between classification efficiency and circulating load, and the impact on the performance and capacity of ball mill circuits. However, the effects on HPGR performance are significantly different. Metso PTI has initiated a pilot scale test program to evaluate the size reduction performance of the HPGR under different circuit configurations. A series of pilot scale HPGR locked cycle tests were performed to simulate the following circuit configurations: including closed circuit with a screen, partial product recycle and edge recycle. This paper discusses the results from the pilot scale HPGR and laboratory scale ball mill test programs. The effect of circulating load and classification efficiency on the performance of ball mill circuits is compared to the effect on HPGR circuits. The fundamentals of grinding behavior are also discussed in order to explain the difference."

Jan 1, 2014