Search Documents

By Robert Dunne

A review of the current ?state of the art? of refractory gold treatment will be provided. This will include a global perspective of the refractory processes that are being practiced, where they are and some discussion on why the processes were selected. Some discussion on the challenges that face the refractory gold industry will also be provided.

Sep 1, 2012

By V. Dementiev

Pyrite cinders is a complex raw material, in which the valuable metals are: Au(2.04 g/t), Ag(35.0 g/t), Cu(0.38%), Zn(1.1%) and Fe(56.0%). In this case high costs for mining and ore processing are not required. The hydrometallurgical treatment of pyrite cinders using thiourea, sulfite and cyanide solutions are examined in this paper. It was found that thiourea leaching is ineffective due to high consumption of thiourea and sulfuric acid. The cyanidation on pyrite cinders was conducted. Various options of the preparation of the products for the subsequent cyanidation have been examined. The tests on cyanidation process optimization were conducted. The effectiveness of these cementation methods by metallic iron or extraction for copper recovery from water phase has been shown. Zinc precipitation can be found in the form of carbonate. In order to recover noble metals from cyanide solutions, the cementation onto zinc dust was tested. The developed technology provides for recovery at 49-56%Au, 20-25%Ag, 22.5%Cu and 50%Zn during NaCN (1.4 kg/t) and CaO (6 kg/t) consumption. The results of pilot plant testing are the basis for technical regulations in order to design a pyrite cinders treatment plant. The technology has been examined in pilot-plant conditions. In this case the recovery of 54.5%Au, 45.3%Ag, 30%Cu and 49.9%Zn was achieved to provide saleable products. Keywords: pyrite cinders, thiourea leaching

Sep 1, 2012

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 N. ??dyanik, Z. Kugukaldinova, O. Mishurina, Y. Glazkova, E. Mullina

"This paper describes the process flow of a selective processing method for acidic mine waters formed in the copper sulfide industry and characterized by a high content of copper, iron, and manganese ions. The proposed method comprises a series of sequential processes: the 1st stage is the recovery of Cu2+ ions by precipitation, the 2nd stage is the recovery of Fe2+ and Fe3+ ions as a dispersed phase by acid-base precipitation, followed by recovery in the 3rd stage of Mn2+ ions by converting Mn2+ with an oxidizing agent into insoluble forms of manganese and subsequent recovery of the dispersed phase by electroflotation. The developed method was tested on the acidic tailing storage facility (TSF) water of a mining project in the Urals region. The practical results of the implementation of the developed method have shown that its implementation will allow to recover copper, iron, and manganese in the form of salable products with a simultaneous decrease in their concentrations in adjacent water bodies down to the applicable MPCs. The process is a simple one, does not require any sophisticated expensive equipment and can be implemented either as a standalone solution or as part of existing treatment facilities. INTRODUCTION As a result of unsustainable utilization of natural resources in Russia, only a small fraction is processed into commodity products, while the rest is systematically dumped as emissions, effluents, and other wastes, polluting the environment. By amount of waste and concentration of valuable components, those wastes can be classified as aqueous technogenic resources [1, 2]. Therefore, the development of processing methods for this mineral resource category is an urgent and priority task that will significantly improve the industry’s profitability and the environmental conditions in the respective regions. MATERIALS AND METHODS Following research methods were mainly applied in this study: lab-scale and pilot-scale tests in a diaphragmless two-chamber electric flotation cell; for solutions and recoverable process products — chemical analysis using photometric determination techniques for metal cations and atomic adsorption determination techniques for the processing products. The phase composition of the precipitates formed in the process was investigated by X-ray phase analysis in the general-purpose diffractometer with a copper anode DRON-1."

Jan 1, 2018

By M. Amani, A. Akbarinasab, M. Bazmandeh

The fourth iron ore production plant at Golgohar mining and industrial company consists a ball mill and wet medium intensity magnetic separators (3500 Guss). The ball mill product applies directly into MIMS and the MIMS concentrate directly transfers to the low intensity magnet separators (LIMS). The thickener is used for dewatering of MIMS and LIMS tailings. By sampling from streams of ball mill and separators, it was defined that the grade and iron recovery were 60% and 80% respectively while they would be 66% and 90% respectively based on the designing documentation. The average of P80 in the ball mill was 342 micron in a month that is coarser than optimized value. This problem was related to unsuitable operation of grinding. For optimization of grinding efficiency, the amount of charge volume was investigated as a variable. The measurements illustrated 32% charge volume. For increasing grinding efficiency, the charge volume was increased 1% and 1.6% in two stages. The results showed ball mill power draw average reached from 3650 kw to 4120 kw and the average P80 reached to 329 micron. All of parameters calculated about 35 day before and after change. Average of Ball mill feed iron ore grade before and after change was 49.24% and 50.32% respectively. This change caused an increase in efficiency of MIMS, where the enrichment ratio and concentrate grade increased 2% and 1.9% respectively. For increasing the iron grade of the MIMS concentrate; the orifice of the tail flow exit pipe changed from 4 inch to 6 inch and the concentrate grade increased 2%.

Jan 1, 2014

By I. A. Makalin, V. N. Yakovlev, S. V. Imangulov, I. V. Zyryanov

"The presented studies aimed at solving the problem of recovery of low-luminescent diamonds, which are not recovered using X-ray luminescence method of processing of diamond-bearing ore. The purpose of the studies is the determination of possibility and feasibility of recovery of such crystals using dual-energy X-ray transmission separation technology. In preparation for studies methods were developed for testing technical and process parameters of XRT-400 Aliud GmbH (Germany) sorter, manufactured according to design specifications of Yakutniproalmaz Institute, and XRT/D TOMRA Sorting Solutions GmbH (Germany) sorter, designed for processing of -50+5 mm ore, as well as RGS-1M, RGS-2 X-ray sorters, developed by JSC NPP Bourevestnik (St. Petersburg) in cooperation with Yakutniproalmaz Institute, for material size of -6+3 mm. In 2018, in-process tests of RGS-OD-3 X-ray sorter prototype on material size of -3+1 mm are scheduled. Studies to determine the content of low-luminescent diamonds and feasibility of their recovery using XRT-400 and XRT/D sorters at the Processing Plant #3 of Mirny Mining and Processing Division (MPD) were carried out. The results of the studies confirmed presence of such diamonds and possibility of their recovery using the dual-energy X-ray transmission method. In-process tests of RGS-2 X-ray graphic sorter were carried out at the Processing Plant #16 of Nyurba MPD. According to the results of sampling of RGS-2 sorter separation products, the recovery of -6+3 mm sized diamonds from tailings of X-ray luminescence sorters was achieved at the level of 96%, while the diamonds recovery was 98.5% in value terms, the concentrate yield was 0.1%. In-process tests of RGS-1M sorters at final recovery of the grease separation concentrate at the Processing Plant #3 of Mirny Division were carried out. Recovery of low-luminescent -6+2 mm sized diamonds by RGS-1M sorter amounted to 92.4% for single operation, and with the double control operation reached 99%. As a result, -6+2 mm sized base product recovery increased by 0.2%."

Jan 1, 2018

By M. Schneider, J. Regler, Th. Neesse, AKW A. + V GmbH & Co., J. Dueck, F. Donhauser

"In practice the advantages of high-quality hydrocyclone separation are often reduced due to varying feed suspension properties, solids concentrations and particle size distributions. Therefore, a new control system for a battery of hydrocyclones with a nominal diameter of 150 mm was developed for bentonite slurry regeneration associated with tunnel excavation using bentonite slurry as a transport aid and lubricant.As can be seen from Fig. 1, the hydrocyclone battery has been modified for regulation purpose. The battery was manufactured with an overflow collector as a closed pressure chamber (1) with a discharge pipe (2) where the throttling valve (3) was installed.With increased stepwise throttling of the overflow, the volume split is changed with the effect that the solids discharge in the underflow is increased. To stabilize the total throughput, the feed pump speed is simultaneously increased so that the pressure drop ?p across the hydrocyclones (9) remains approximately constant. However, this causes the increasing pressure inside the hydrocyclones, which results in a further intensification of the underflow discharge. After the detection of the break through of the air core (transition state) by the capacitive sensor (4), the throttle valve is shortly opened, and a new regulation interval can start again. The control concept is focused on maintaining the optimum operating state, characterized by the underflow discharge shape at the transition state between rope and umbrella discharge."

Jan 1, 2003

By K. Hori

In this study, the determination of optimal conditions for jig separation has been investigated with a TACUB jig separator. The TACUB jig, from which a Batac jig originated, is a useful gravity separator that is effective to separate crushed solid wastes as well as coarse ores. A mixture of crushed polyethylene (specific gravity 1.3) and polyvinyl chloride (specific gravity 1.5) was separated by a batch TACUB jig by varying the frequency (wavelength) and the amplitude (pulse stroke length). It was found that the sample particle layer in the separation chamber was fluidized without turbulence by the rising water pulse resulting in the separation. However, a too strong water pulse reduced the efficiency of the separation of the particles. The results indicate that the specific velocity of the rising water is necessary to fluidize the particle layer without turbulence to separate the particles. The expansion ratios of the thickness of the particle layer in rising flows were measured in a cylindrical column. The pressure losses by the particle layer were calculated from the expansion ratios and the velocity of the rising flows by Ergun?s equation. The curve of the pressure loss vs. the velocity of rising flow showed a maximum point corresponded to the point of fluidization. The velocity of a rising flow on the point of fluidization provided the optimal separation condition.

Jan 1, 2014

By Rod Hocking, Nenad Djordjevic, Randolph A. Pax

"Semi Autogenous (SAG) mills exhibit complex interactions between the competing processes occurring inside them. Traditional practice is to measure the operation of the SAG mill at an operating point in terms of its load, power draw, steel load, ore throughput rate, water addition rate and more recently the feed size distribution. However significant departures from standard operating conditions can occur due to changes, for example, in ore hardness, the quest for higher throughputs and mill internal geometry.Fundamental to SAG mill operation are the internal processes that ensure that a given ore throughput and discharge grind size are obtained and maintained. These processes can be measured using non-contact acoustic emissions from the SAG mills which utilise the information rich noise that is generated by the grinding, flow and charge movement inside a mill. Non-contact acoustic in-mill measurements can be used for the validation of discrete element modelling (DEM). The approach being reported in this paper is to model the individual physical processes and then to simulate the assemblage of a physical mill.By modelling the assemblage of steel balls in a 1 m pilot mill and duplicating the experimental arrangement exactly with the PFC3D model, it was possible to replicate the pattern of falling balls in the experimental mill. Furthermore, the key features of the acoustic emission spectrum that was calculated compared favourably to the measured spectrum at low frequencies. The experimental spectrum is shown in Figure 1."

Jan 1, 2003

By Sebastian Fredes, Williams Calderon, Mauricio Toledo, Gonzalo Montes-Atenas

The influence of the temperature on leaching kinetics of copper-bearing minerals has been widely acknowledged: the leaching of copper oxide ores is generally described as exothermic reactions while copper sulfide ores are associated with endothermic reactions. Reports studying the influence of temperature on leaching kinetics at laboratory scale frequently involve the implementation of either fully flooded isothermal systems or jacketed column setups. Nevertheless, at industrial scale controlling or monitoring the temperature or temperature changes is almost impractical. In fact, leaching process is usually a large tonnage operation and changing the temperature of vast amounts of material would require an enormous amount of energy and highly specific heat transfer equipment. Preliminary computations on the amount of energy required to increase the temperature of a heap or dump in 1°C indicates that it would be impossible to find an energy source able to accomplish the task. In order to provide a solution to this matter it is suggested to create an energy managing protocol (EMP) that permits heat up or down the porous body locally taking advantage from the temperature gradients and from the reactivity of different regions within the material as a function of the leaching cycles. The aim of this study is to evaluate different heat transfer regimes in a 1 ton pilot column leach operation. It was found that best results of copper recovery are obtained when heating up the column in horizontal bands starting from the top of the column moving downwards as the leach cycle evolves. The system is expected to be suitable for industrial applications, but further development of the technique is required, particularly in the design of the heat transfer dispositive.

Jan 1, 2014

By James A. Finch, Linda Restrepo, Mitra Mirnezami, Amy Buchanan, Maryam Hashemi

Settling rate studies on slurries from Cu/Zn flotation plants occasionally reveals two aggregation maxima at pH 8-10 and 10-12 (Figure 1). From measurements on sphalerite and silica suspensions the double maxima could be replicated. These maxima correspond to the effects of Zn and Mg ions. The aggregation mechanism of both ions is related to the formation of the corresponding hydroxide precipitate. A combination of chemical and electrostatic bridging by the hydroxides is considered to be the aggregation mechanism.

Jan 1, 2003

By F. O. Milovich, V. A. Ignatkina, L. A. Samatova, E. D. Shepeta

"The aim of the research was to increase technological parameters of the flotation of finely disseminated scheelitecarbonate ore using selective reagent regime.In the studies the monomineral fractions of scheelite and calcite; samples of the scheelite-carbonate ore of the Primorsky deposit with the range of the carbonate module 3-65 were used. A carbonate module of ore is the ratio an initial assay of calcite to an initial assay of scheelite. Sodium oleate, iso-decyl alcohol, water glass, calcium chloride, soda or sodium hydroxide as flotation reagents were used. The flotation method in the Hallimond tube and in a mechanical cell, an adsorption method for study the kinetics of adsorption of oleate and isodecyl alcohol and its combinations on scheelite and calcite, IR spectroscopy including ATR to study the composition of surfacecompounds were applied. Selective reagent regime may include two ways - a combination of collectors or a modification of depressants of calcite and other gangue minerals. On one hand the combination of sodium oleate and higher fatty iso-alcohols of fractions C12-C16 in the rough scheelite flotation to reduce the flotation activity of calcite has been studied. On the other hand depression of calcite flotation by calcium chloride with water glass into rough flotation was used; modified the steaming cycle by Petrov's method applying change the concentration of water glass and other. Fatty iso-alcohols can adsorb on the surface of the calcite that reducing the adsorption of sodium oleate and floatability of calcite has been fallen. Their co-adsorbing with oleat on the scheelite surface were support the hydrophobization and flocculation of slime particles of scheelite. Flocculation of the particles of scheelite leads to increased up 3 times of the recovery of scheelite from the slime class, and calcite change up only 1.4-1.1 times it’s recovery.The growth of the carbonate module more than 11 is decreased the assay of WO3 in the scheelite concentrate of selection cycle."

Jan 1, 2018

By Gülay Bulut, Ferihan Göktepe

Pyrite is the most abundant and widespread sulphide mineral and its flotation has been investigated by many researchers. It is well known that, pyrite is floatable under acidic conditions and recovery drops with increase in alkalinity due to changes in solubility of iron hydroxides on the pyrite surface. Most of the works also revealed out that pyrite displays variable behaviour in flotation circuits. This behaviour mainly related with origin and mineralogical composition of ore, liberation size, electrochemical behavior of minerals constituting ore, the trace and minor element content of the mineral, texture, application variables belonging to flotation circuits. In this study, copper activation effect was examined on different origin of pyrites. The experiments were performed with five high grade samples of pyrite from the different mining districts of Turkey. The pyrite samples were different from each other with respect to either their physical or chemical properties; they were different physical appearances and contained different associated minerals. In the micro-flotation tests, 1x10-4 mole/L copper sulphate and 1x10-4 potassium ethyl xanthate were used as flotation reagents. First of all pyrite samples were floated using only xanthate at different pH. Results showed that best floatation results obtained at acidic pH range for all pyrite samples. When copper ions were used to activate, all pyrite samples floated at neutral and basic pH values, too but depending on the samples, recovery and pH ranges were slightly changed.

Jan 1, 2014

By X. L. Song

In this paper, ZrSiO4-Cd(S1-xSex) inclusion pigments were synthesized by hydrothermal method. Microstructure, morphology and properties of the powder were characterized by X-ray diffraction (XRD), transmission electron microscopy (TEM), infrared spectroscopy (IR) and CIE chromaticity, respectively. It shows that, precursors obtained at pH is about 6.0 are sol and products with higher crystallinity, red hues (a*=38.23) and inclusion ratio when hydrothermal synthesis temperature is 180? and holding time is 12h. Hydrothermal reaction mechanism was further analyzed as well. In particular, an ideal model of pigment formation was proposed.

Jan 1, 2014

By Ayumu Tokumaru, Shuji Owada, Gangfeng Liu

"Several million spent catalytic converters are disposed annually worldwideandeach containing about 0.5-1g ofplatinum and 1-2g of all the platinum group metals (PGMs), then,it is fairly significant to recover PGMs from the spent automobile catalyst. The PGMs’ recovery is a proven technology, including pyro-and hydrometallurgicalmethods, and the highestrecovery can achieve approx. 98 wt.% by acid-leaching method, but theserecyclingtechnology seems to be too highlyenergy-consuming. SincePGMs aredistributed only in coat layer in spent catalyst, it is of great significance to enrich the coat layerwith some physical methods as a pretreatment.In this paper, according to the double-deck structureof the catalyst andthe different properties of the two main compositions, alumina-ceria phase of coat layer and cordieritephase of substrate, the concentration of the coat layer was investigated by combiningthree kinds of physical separation technologies, heating-quenching, crushing and surface grinding. In the first stage, thermal stresscouldgeneratebetween the two phasesby thermal expansion difference,to producefracturesin the thinnest part of coat layerand micro-cracksin coat layerand at the interface of coat layerand cordierite substrate. The second stage of crushing, by jaw and roll crushers, could break dominantly the coat layer and the third stage of surface grinding, carried out by attritor, promoted further separation and concentration of coat layer fromthe substrate.We also improved the attritor grinding by adding a new lift-arm to the cell for achieving better mobility of sample particles in the selective surface grinding."

Jan 1, 2014

By V. Panayotov

Flotation practice usually uses the Sheridan-Grizvold selection method, where zinc minerals are activated by addition of copper sulphate and depressed by addition of sodium cyanide, ferrous sulphate, zinc sulphate or various mixtures of these reagents. The method is suitable for the flotation of sulphide ores, which, do not bear precious metals, such as silver and gold. When classical technology with cyanide is applied to depress zinc minerals, the precious metals, which are contained mainly into lead minerals, are dissolved into the solution, which represents a technological loss. In order to avoid this process, we have made use from the knowledge on the semiconducting properties of lead and zinc sulphide minerals. We have changed the surface charge of minerals by means of an electrochemical impact realized by direct current application and specially designed set of insoluble or semi-soluble electrodes. Thus, flotation separation of lead and zinc sulphide minerals was achieved, at increased recovery of lead, zinc and silver. The technology proposed was implemented in a Bulgarian flotation factory; it worked over 5 years providing significant, positive economic, technological and environmental impact. Silver recovery was increased by 0.8-1.0 %, and lead and zinc recovery was increased by 0.8-1.0 %. The technology is easily controllable and adaptable to raw materials and parameters of the technological process. It can be paid back for 6-8 months after the implementation. Keywords: lead-zinc flotation separation, electrochemical impact

Sep 1, 2012

By Victor Avdokhin, Valery Morozov, Valery Stolyarov

"The adaptation-determination method provides for keeping given relationship between concentrations of ionic and molecular components of a pulp liquid and correction of parameter of the model on the basis of analysis of ore composition, technological parameters of flotation, and cost of valuable components.• optimisation program “ADDEFLOT” and combined analyzer of pulp “RA-931”An adaptation-determination method provides for two-level control of bulk and selective flotation of polymetallic ores. The lower level provides for control of reagent dosing on the base of control of ionic composition parameters in the pulp liquid phase. The control goal is to keep the basis relationship among parameters of ionic composition (ions A, B, C) in pulp liquid phase given by equations of type:• alg[A] + blg[B] + clg[C] = K (1)Values of a,b,c coefficients are uniquely determined by a physico-chemical model used. Kind of the model depends on mineral composition of an ore and reagent regime applied. As parameter of ionic composition, also values of one or two parameters (or their combinations) can be used, for instance, pH, concentration of sulfide ions, ions of collector.The upper level of control is based on calculation and use of technological results of selective flotation. At the level, an adaptation of the model (1) is conducted. A principle of the adaptation is in determining (and including in pre-setting function of the lower level) such value of K parameter, at which maximal economic effectiveness of the process is reached. As a criterion of economic effectiveness of the process, a function of losses of several valuable components, taking into account cost of the metals lost, is used. For 2-component ore, in simplified form, the function is as follows:"

Jan 1, 2003

By Eduard Stefanescu, James M. Mwase, Egon Sehner, Stoyan Gaydardzhiev

"The study is focused on investigating the production of synthetic rutile of 95% purity or higher from a Norwegian ilmenite ore concentrate. This was to be achieved using a sequential process involving oxidative-reductive roasting at 975°C, followed by magnetic separation to further concentrate the sample and finally leaching with 20% HCl to remove mostly Fe-based impurities leaving behind synthetic rutile. The key process parameters considered were roasting time (1 and 2 hrs), leaching temperature (100, 125 and 145°C), pulp density (15 and 25wt%) and leaching time (3 and 6 hrs). The option of directly leaching the concentrate without pre-treatment was also evaluated. Test work was conducted using samples screened to a size fraction of 105-212 µm. Samples were leached in a 1 L glass autoclave. The reactor was hermetically sealed and thus the resulting pressure was not controlled but was dependent on the process temperature. The product from leaching was calcined and analysed via XRD and SEM to determine the purity of the product. The magnetic separation step was found to be ineffective and discontinued early in the study. Thus far a product with 97% synthetic rutile purity as analysed by SEM has been produced by directly leaching the concentrate at a temperature of 145°C, pulp density of 25wt% and leaching time of 6 hrs.INTRODUCTION The study investigates the use of a process route comprising oxidative-reductive roasting, followed by magnetic separation and then hydrochloric acid leaching and calcination to produce synthetic rutile from ilmenite concentrate. The grade of synthetic rutile produced should be suitable for chlorination to produce pigment grade TiO2. Hence the proposed process should produce a product consisting of no less than 95% synthetic rutile. Additionally, the product should have no more than 1.5wt% combined content of CaO and MgO (Zhang et al., 2011) as they tend to form high-boiling-point chloride leading to the malfunctioning of fluidized bed chlorinator (Cardarelli, 2008). Finally the product should have no more than 2.1wt% of Fe2O3."

Jan 1, 2018

By Przybycien Krzysztof, Tumidajski Tadeusz

"Materials being the feed to mineral dressing plants are the mixture of ore minerals having different physicochemical properties among others, especially susceptibility on crumbling. During exploitation of material and their preparation to communition the big number of narrow classes comes into being and they have different properties. The determination of common communition laws of multicomponent (on account of size and mineralogy) materials leads to the estimation of mutual influence of narrow classes and their rates, taking into account their endurance and mineral density, on general ratios and their simultaneous crumbling.If we try to accept the grain membership to the certain component as the differing feature, we can write the balance equation as: ( ) ( ) S= = k i i i g x p f x1where: fi – density functions of one type distribution and of different statistical moments (characterising the homogeneous materials grinding), pi – “stirring” proportions of each components, S== k i i p 1 1 . From the point of verification view of detailed versions of this equation the inverse operation should be done: in the base of the components’ mixture grinding results and the results of components grinding the legitimacy of equation should be checked.The using of finished distribution mixtures’ theories to the grinding process analysis allows, first of all, to verify the assumption of grinding independence of individual feed components. Talking more generally it would lead to the comparison of grain size distributions, obtained by individual components grinding from the defined in the mixture distributions theory calculations. One of experimental and calculation results for artificial mixtures of galena and dolomite is presented on figure below (df means distribution function)"

Jan 1, 2003

By E. V. Manlapig, M. C. Harris

"The complexity of the flotation process has made the identification of an appropriate model structure for the predictive simulation of flotation circuits far from obvious. Work performed under the auspices of the AMIRA P9 flotation program has indicated that the metallurgical performance achieved from a flotation cell arises from the interaction of two essentially independent property classes, namely the cell operating conditions and the properties of the cell feed stream. These observations have led to the development of a model structure that associates a constant set of flotation parameters to the streams of the circuit, which are transformed into a flotation response depending on the operating conditions within the flotation cell.A variety of alternative model structures exist within the literature which do not make this distinction between cell and stream flotation properties, and many of these models are in widespread use for flotation optimisation and design studies, both in industry and academia.In this paper, circuit survey and laboratory batch flotation data collected from an industrial flotation circuit are analysed. The data demonstrate that streams in an industrial flotation circuit exhibit varying particle floatability – concentrate streams typically float faster than tailing streams (Figure 1(a)). This floatability, however, is generally conserved around the various units within the flotation process (i.e. the effective floatability of the combined feed and product streams of a unit is equivalent (Figure 1(b)))."

Jan 1, 2003