Search Documents

By Kevin Galvin, James Dickinson

"The Reflux Flotation Cell (RFC) consists of an inverted fluidized bubbly bed, located above a system of parallel inclined channels. This novel arrangement enhances the hydrodynamics of flotation in a number of ways. Firstly, the free-surface of the flotation cell is enclosed via a fluidized bed distributor, ensuring uniform application of wash water to promote desliming. Secondly, foams can be variable in their stability and lack permeability. Hence this system is preferably operated in the absence of foam, instead opting for a bubbly zone of high bubble volume fraction through the use of a very high gas flux and/or wash water flux. Discharge of the product proceeds upwards through a central port of narrow cross-sectional area. Thirdly, the system of inclined channels increases the segregation between the bubbles and the water, ensuring the bubbly zone remains contained within the upper vertical zone, and the liquid flux reporting to the overflow is low. The expanded operating regime of the RFC was investigated by covering a very broad range of gas fluxes, up to 5.5 cm/s, and wash water fluxes, over 2 cm/s. Both laboratory and pilot plant performances were quantified by the recovery of hydrophobic particles and rejection of the hydrophilic particles, using model and industrial coal feeds. The new system hydrodynamics was found to offer significant advantages in flotation kinetics, desliming, and particle recovery across a broad particle size range. Bubble interfacial flux reached as high as 600 m2/m-2s-1. The upper system consists entirely of a high volume fraction of bubbles, which provides a "safety-net" to protect against the detachment of coarse particles, and a highly permeable zone for counter-current washing of the flotation product. Under relatively extreme conditions selective stripping of the product is observed, resulting in grade recovery values beyond those achievable using the tree flotation method."

Jan 1, 2014

By Sanjay Chaudhuri, V. K. Kalyani, Pradeep Kumar Singh, T. Gouri Charan

"Indian coal industry is the world's third largest in terms of production and fourth largest in terms of proven coal resources. Indian coal deposits are generally of high ash content varying from 24 to 50%. The coking coal reserves of India constitute about 12% of the total coal inventory, and the coals are washed in washeries, before being used for metallurgical purposes. The processes used for washing coking coal is Jigs, Heavy medium bath and Cyclones for the sizes upto 0.5mm, while the coal fines are washed in spirals and conventional flotation cells. Extensive studies were carried out for understanding the performance of an operating conventional flotation plant. The performance studies revealed that appreciable quantity of clean coal is being lost in the tailings due to varied reasons and the required quantity and quality was not achieved. In this study, laboratory Flotation experiments were carried out to examine the cleaning characteristics of the said tailing and to determine the effective process parameters for recovery of clean coal at desired ash level ( 16-17%). Response Surface Methodology was applied to investigate the operating conditions of the Froth Flotation for modeling and optimization. Primary study reflects that there is a possibility to recover clean coal at 16-17% ash level with a yield of about 60-70% from that Flotation tailing of feed ash about 28%.INTRODUCTIONMost of the Indian coking coals are difficult in cleaning characteristics and high ash content due to its drift origin and its age of formation for more than 225 million years. Indian coking coal reserve is limited about 12% of total coal reserve. The minerals are homogeneously mixed with the coal macerals. The open cast mining practice and mechanized mining methods are being adopted to meet the increasing demands of coals resulting in generation of more coal fines and subsequently the quality is deteriorated. Indian good quality coking coal reserves of higher seams are fast depleting and hence coking coals of difficult cleaning characteristics of lower seams (LVC) are being now explored. Due to increased mechanization of coal mines to cope up the increased demand and followed by size reduction of lower seam coals in search of better liberation , the coal fines are increasing by leaps and bounds. The processes used for washing coking coal is Jigs, Heavy medium bath and Cyclones for the sizes upto 0.5mm, while the coal fines are washed in spirals and conventional flotation cells."

Jan 1, 2018

By S. Sarioglan

The bleaching of edible oils, industrial oils, and some chemicals may be achieved using natural and acid-activated clays. These materials remove impurities and pigments from the oils. In this work, a bentonite clay from Turkey (Somas bentonite) was activated with different concentration of H2SO4 and HCl at 70~90°C and 5~8 hours. Under these conditions several cations were removed from natural bentonite (Al, Mg, and Fe) and DTA-TGA, FT-IR curves of the solids obtained after acid treatment were examined. Treatment time and acid concentration increased the degree of destruction of the bentonite structure which was determined by ICP. Activated samples were tested in order to verify their capacity to bleach soybean oil and compared to standard commercial bleaching clays. The experimental results indicate that treated samples with H2SO4 have the same efficient in bleaching crude soybean oil with the standards and treated sample with HCl is more efficient in bleaching than the commercial products.

Jan 1, 2014

"The first 90 years of the last century generated a considerable volume of technical literature on the subject of grinding mill power and its prediction. It is therefore surprising that during this time there were no published models for predicting grinding mill power draw which were validated using a wide range of comprehensive industrial scale data.Notwithstanding this lack of data, the majority of models placed limited emphasis on the internal dynamics of mills and relied on simple assumptions which considered the charge to take up a fixed position and shape.Laboratory based studies (Liddell, 1986), and industrial data (Morrell, 1993), showed that these assumptions did not hold over a wide range of operating conditions, and brought into doubt the ability of the (then) existing models to accurately predict grinding mill power draw. Furthermore no model treated AG,SAG and ball mills as a single class of devices whose power draw could be predicted with the same equations.Through a detailed investigation of the charge dynamics, coupled with a large data base of industrial mill power draws, this situation was rectified by the end of the century, by which time a model had been developed and validated, which accurately predicted the power draw of all wet tumbling mills. This paper describes this model and reviews some of the earlier attempts at mill power modelling."

Jan 1, 2003

By T. V. Vijaya Kumar

Most of the iron ore washing plants set up in India consist of sizing of the ore by dry / wet screening, washing and classification by screw classifiers followed by multi-stage hydrocycloning of screw classifiers? overflow. The underflow of the hydrocyclones forms the concentrate which is used for pellets making. As the cut-point of hydrocyclones is generally at 20 microns, relatively lower diameter cyclones in a cluster with parallel feeding are used. This is found to, often, result in choking of the spigots, at times, by extraneous material as well reporting along with the slurry. It is observed to lead to sub-optimum performance of the cyclones with loss of iron values into cyclone overflow and thereafter into tailings. Analysis of data of an operating beneficiation plant (Beneficiation Plant - 1, BP - 1) of M/s Jindal South West (JSW) Steel Limited, Karnataka, India over a period of one year indicated that weight recovery of the concentrate from two-stage hydrocycloning was 45% (with respect to feed to the cyclones) and assaying 63.43% Fe, 4.43% SiO2 and 2.32% Al2O3. CSIR - National Metallurgical Laboratory - Madras Centre (NML - MC), India developed flotation process for processing screw classifier overflow with a view to improve the recovery of ultrafine iron values from it. Detailed investigations at laboratory scale flotation column and on-site pilot plant trials proved that the production of the concentrate and productivity of the plant could be enhanced by adopting reverse cationic column flotation technology. It was established that a typical screw classifier overflow analyzing 60.54% Fe, 5.95% SiO2 and 3.67% Al2O3 could be upgraded to 64.21% Fe, 2.64% SiO2 and 2.20% Al2O3 with 82.7% weight and 84.7% Fe recovery. This translates into an increase of 80% in weight recovery of the concentrate compared to what is being obtained, at present, by two-stage cycloning. It had an added advantage of reduced SiO2 in the concentrate. Thus, SiO2 in the final concentrate could be reduced to 2.64% from 4.43% by adopting column flotation. The concentrate so generated meets the stringent specifications, especially in terms of alumina in it(should be less than 2.50%) for pellets making. Later, JSW Steel Limited set up the flotation plant in BP - 1 based on the process developed by NML - MC and was commissioned in 2010. Keywords: iron ore, washing, ultra fines, alumina reduction, column flotation

Sep 1, 2012

By A. V. Neretin

This paper is a continuation and extention of the previous paper on energy-efficient placer mining methods which was presented in 2007 at the World Gold Conference in Cairns, Australia (Nere-tin, 2007a). The paper reports the results only on gravel processing as they were not fully discussed in the previous paper. It is common knowledge that placer gold mining requires processing of large tonnages of gold-bearing gravels by wash plants of different types. In Russia, gold-bearing gravels are mostly disintegrated and screened by the following types of wash plants: a water canon, a derocker or a scrubber trommel. Gold-bearing gravel processing is regarded as highly energy-intensive due to high energy consumption for transportation of large tonnage of material and high pressure water flow rate. The industries requiring a significant expenditure of energy normally estimate the process efficiency by energy consumption per product unit. In order to obtain quantitative data on energy consumption for gravel processing and to estimate placer mining costs, in 2003~2006 four placer gold mines located in the Irkutsk Region were thoroughly investigated through time and motion studies. At all these placer mines gravels were processed by a water canon, a derocker or a scrubber trommel. The objective of the study was to investigate the following parameters: ? wash plant throughput rate per hour; ? hourly diesel fuel and power consumption per cubic meter of gravel processed; ? volume per cent and shape of oversize material; ? different combinations of gravel processing methods. The results from this study were as follows: ? specific energy consumption per cubic metre of gravel processed for different combinations of gravel processing methods was estimated; ? a correlation between energy consumption and unit cost of gravel processing was determined; ? effect of yield and shape of oversize material on wash plant throughput rate per hour was de-fined; ? energy-efficient gravel processing methods were developed. Based on these results, an improved method of gold-bearing gravel processing was proposed. Implementation of this method at a placer mine saved US$ 123 000 in operating expenses only during a short-time Siberian placer mining season. At the same time, this method improved gravel disintegration, screening and processing operations used at the mine. Another advantage of the pro-posed method was higher gold recovery due to reduced yield of undersize material. In order to improve efficiency of gravel processing, a wash system including a derocker for primary gravel processing and a lightweight scrubber trommel was developed. There are two economic benefits of replacing water cannons with the proposed wash system. The first benefit includes lower operating costs for gravel processing due to lower specific energy consumption, while the second benefit is lower gold losses during slurrifying (disintegration and screening) and sluicing of the undersize material.

Jan 1, 2014

By Chuanyao Sun, Zhenguo Song, Aijun Ren

"Lead-Zinc sulfide ore containing carbon is one of typical refractory resource. Because organic carbon has natural floatability, it has a great negative influence on flotation of lead-zinc sulfide ores. It is difficult to obtain qualified lead concentrate by flotation using conventional reagent system. In order to reduce the influence of organic carbon on lead and zinc flotation, two alternatives, i.e., removal carbon by pre-flotation or adding carbon depressant in lead and zinc flotation, can be adopted. If the flowsheet of removal carbon by pre-flotation is adopted, the lead and zinc recovery which are lost in the carbon tailings is often higher. If conventional carbon depressants are used in lead and zinc flotation, it has a certain depression effect on lead minerals. In the paper, the ore grading 1.49% Pb, 5.73% Zn, and organic carbon 3.14% has been carried out by flotation tests. In the bulk flotation of carbon and lead minerals, Pb concentrate is only about grading 21.00% Pb. In the pre-flotation of organic carbon, it’s hard to remove most of carbon, and the loss of Pb and Zinc recovery is too much in carbon tailings. New carbon depressant BK547 and three conventional carbon depressants including sodium lignosulphonate, Fe-Cr-lignosulfonate, and dextrin, are used in comparative tests of Pb flotation, and results show that BK547 have a good depression on organic carbon, and have almost no effect on Pb flotation. Using BK547 as carbon depressant, cycle test in flowsheet of two-stage Pb and Zn selective flotation and coarse concentrate regrinding have been carried out, and results show that concentrates grading 62.51% Pb and 41.33% Zn are obtained at recoveries of 70.24% Zn and 83.58% Zn.INTRODUCTION Lead and zinc mineral resources are still the main source of lead and zinc metal. For lead-zinc sulphide ore, lead concentrate and zinc concentrate are obtained by flotation. Minerals separation by flotation is a complex process and is dependent, along with other factors, on the relative distribution of hydrophobic (e.g. collector, polysulfide, elemental sulfide) and hydrophilic (e.g. metal hydroxide) species between different particles of the same mineral phase and between mineral phases (Piantadosi & Smart, 2002; Boulton, Fornasiero & Ralston, 2003). Organic carbon has natural floatability and is often associated with lead-zinc sulphide ores. If the content of organic carbon in the ore is more, the flotation index will become worse. The main problem is that the concentrates has too low grade to be qualified grade products for sale. At the same time, the consumption of reagents is more than ordinary ores without organic carbon. In order to solve the influence of organic carbon on the flotation of lead and zinc and other sulphide ores, a lot of research works have been carried out."

Jan 1, 2018

By Liu Yong

A complex rare metal ore, containing various valuable elements is used in this study. The mineralogy results indicate that the mineral composition of the raw ore is complex, and there are more than thirty-eight kinds of minerals in the raw ore, and the elements, such as Nb, Ti, Fe, REO, Be etc, always form isomorphism relationship. The REO content of the raw ore is 0.45%, and the contents of other rare metals Nb, Ta, Ti and Zr are 0.33%, 0.02%, 0.93% and 2.00% respectively, and the main gangue elements are Si and Al, and their contents are 71.84% and 8.5% respectively. The beneficiation-metallurgy combination process of ?preconcentration-reduction and magnetic separation-acidification and leaching-precipitation? is used to treat the raw ore. The preconcentration of raw ore is realized through high intensity magnetic separation, and the magnetic field intensity is 1.0T, and the rough concentrate with 2.88% REO and 1.95% Nbresidue can be obtained, and the iron recovery is 92.83%, and the REO recovery is 95.87%, and the Nbresidue is treated by hydrometallurgy process of ?acidification-leaching-precipitation?, then the rare earth oxide, niobium concentrate and zircon-bearing residue can be obtained, and the REO recovery is 80.10% and the Nbresidue and the gangue obtained from the high intensity magnetic separation process is separated by gravity separation, and then the zirconium concentrate can be obtained. The indexes of the beneficiation-metallurgy combination process used to treat the complex rare metal ore are as follows. The REO content of the rare earth oxide is 92.58% and the recovery is 58.44%. The Nb2O5 and 27.42% TFe can be obtained. Iron in rough concentrates is removed after reduction and low intensity magnetic separation, then the iron powder with Fe content over 88% and nonmagnetic rare-metal 2O5 recovery is 94.77%. The rare-metal 2O5 recovery is 80.04%. The mixture of the zircon-bearing 2O5 content of niobium concentrate is 28.34% and the recovery is 54.02%. The ZrO2 content of zirconium concentrate is 28.43% and the recovery is 52.35%. The TFe of iron powder is 89.53% and the recovery is 78.13%. Keywords: rare metal ore, separation and extraction, beneficiation-metallurgy combination, comprehensive utilization

Sep 1, 2012

Micro-fine disseminated molybdenite was ground in a laboratory stirred mill using steel media and zirconium media respectively. The floatability of ground molybdenite and slurry chemistry was investigated. It was found that the total recovery of ground molybdenite using zirconium media was higher while the pH and Eh of the slurry are lower than that using steel media. When steel media was used, iron oxy-hydroxides had been generated coated on the surface of molybdenite, and decreased the floatability of molybdenite, especially of the -10µm particles. The iron and lead oxy-hydroxides was generated by the reaction between steel media with pyrite and galena associated with molybdenite. It was also found that changing the grinding media to inert can benefit the flotation of molybdenite. Keywords: stirred mill, grinding environment, floatability, micro-fine disseminated molybdenite, fine grinding

Sep 1, 2012

By M. Saghaeian, R. Hejazi, M. Mohammadinejad, M. Zare

"Water quality and quantity at the Tailing Thickening Section of processing plants is considered to be a crucial factor and impacts the performance of several sections. Therefore, gaining stable and effective performance at the Tailing Thickener Section is one of the principal goals of minerals processing plants. Different factors such as: qualitative and quantitative sustainability of plant feed, the balance between the mass of input and output water flow, the dissolving method and distribution of flocculant and the amount of dirt, solid percent of thickener inlet pulp, affect the performance of Tailing Thickener. The present study investigates the successful improvement of Tailing Thickener performance at Jalal Abad Hematite Iron Ore Concentrate plant through optimizing the above operational parameters. For this purpose, the following optimizations have been done: homogeneity and sustainability of plant input feed rates, fine pulp removal collected by the bag filters at the input of the thickener, adding an quietness box before moving the tail pulp into the thickener to reduce the pulp turbulence, replacing flocculant pump and injection pipes, improving water mass balance which leads to optimizing and stabilizing the concentration of input tail pulp into the thickener, adding a mixer to increase the flocculant preparation time, increasing the locations for adding flocculant to thickener from 1 point to 6 points. Applying the above modifications and monitoring the Tailing Thickener performance over the course of months, has led to the improvement of recovered water quality from 500 ppm to 200 ppm, increasing the solid underflow of the thickener from 40% to 48.5%, and similarly, reducing water loss by 60%. Of other important effects, due to improvement of water quality, grade of final concentrate 0.5% increased and because of 73% increase in monthly working hour, concentrate tonnage 55% increased. Finally, sludge thickener optimization has had a significant effect in selecting and designing the dewatering system from the underflow of the sludge thickener."

Jan 1, 2018

By M. Abdollahy

Solvent extraction of gallium from Bayer solution of Jajarm Alumina plant was investigated us-ing Kelex100 as extractant. Taguchi experimental design method was used in order to optimize the separation factor of Ga and Al. Optimum conditions were obtained as follows: Temperatuer: 65°C, volume ratio of aqueous phase to organic phase: 3.0:1.0, process time: 60min and organic phase containing 15vol.% Kelex100, 5vol.% isodecanol in kerosene. In this conditions, 74% gallium and 0.3% aluminum was extracted and separation factor was obtained 828.63.

Jan 1, 2014

"The first 90 years of the last century generated a considerable volume of technical literature on the subject of grinding mill power and its prediction. It is therefore surprising that during this time there were no published models for predicting grinding mill power draw which were validated using a wide range of comprehensive industrial scale data.Notwithstanding this lack of data, the majority of models placed limited emphasis on the internal dynamics of mills and relied on simple assumptions which considered the charge to take up a fixed position and shape.Laboratory based studies (Liddell, 1986), and industrial data (Morrell, 1993), showed that these assumptions did not hold over a wide range of operating conditions, and brought into doubt the ability of the (then) existing models to accurately predict grinding mill power draw. Furthermore no model treated AG,SAG and ball mills as a single class of devices whose power draw could be predicted with the same equations.Through a detailed investigation of the charge dynamics, coupled with a large data base of industrial mill power draws, this situation was rectified by the end of the century, by which time a model had been developed and validated, which accurately predicted the power draw of all wet tumbling mills. This paper describes this model and reviews some of the earlier attempts at mill power modelling.INTRODUCTIONThe early 1900?s saw the beginnings of published papers which tackled the problem of what influenced the power draw of tumbling mills and what equations (models) were suitable for predicting power draw. This interest continued for the rest of the 20th century and the literature contains many papers that attempted to satisfactorily answer these questions. A common and disappointing theme in the majority of them, however, is the distinct lack of suitable experimentally derived power draw data. This deficiency consequently limited the practical application of many of the attempts at modelling mill power draw. Prior to the early 1990?s some limited data on the power draw of grinding mills in the literature can be found, but as Harris et al (1985) remarked they are ?... too frequently unusable simply because one or more essential variables have been omitted?. The absence of published results, from vigorous experimental testing of the various models, has resulted in a general lack of evaluation of the validity of assumptions and hypotheses that such models contain. Harris et al (1985) however, did make some efforts in attempting to evaluate the performance of various power prediction equations and in so doing further developed a semi-empirical one of their own. However, in using manufacturers? published data, which themselves were undoubtedly generated by proprietary equations of unproven validity, they probably succeeded in generating little more than a consensus of the various errors in these models."

Jan 1, 2003

By L. Pan

As an air bubble approaches a mineral surface, a thin liquid film (TLF) is formed in between. The excess pressure (p) in the film determines the kinetics of drainage, which is of critical importance in flotation. We have developed a new method of monitoring the kinetics for both stable and unstable wetting films and analyzed the results using the Reynolds lubrication theory. It has been found that initially a TLF is thinned by the Laplace pressure (pcur) due to changes in bubble curvature and subsequently by the disjoining pressure (?) created by the surface forces present in a wetting film. In the present work, we have derived an analytical expression for determining ? from the temporal and spatial profiles of TLFs on mineral surfaces. The results show that ? > 0 in a film formed on a hydrophilic surface and ? < 0 in a film formed on a surface hydrophobized by collector. Thus, the role of collector in flotation is to create a hydrophobic force and hence a negative disjoining present in wetting films. It is also found that smaller bubbles can create higher Laplace pressure, which helps the TLF drain faster.

Sep 1, 2012

By B. Peszko

There are many measuring methods of grain sizes. Taking into consideration the fine grained materials, one of the existing techniques is measurement of grain size by the Coulter Counter method. This method is based on the measuring the changes of electrolyte resistance during grains flow through the calibrated diaphragm. This method is relatively simple to apply, significantly precise and the results are given quickly. However, the problem is that the measurements are limited by the size of diaphragm. Because of this situation, measurement of grain sizes by this instrument is possible in range from 10% to 60% of diaphragm diameter. This limitation causes the necessity of re-placing diaphragm in case when the measured grain size distribution concerns the material of large size range. In purpose of determining the total grain size distribution, the synthesis of some partial distributions, given for individual grain fractions (with different diaphragm applied), is required. This is possible to do by calculating the mixture of distribution functions, created from the partial distribution functions. Two materials were selected to research: quartzite and glass, which were comminuted in cylindrical crusher, additionally grinded, screened and then analyzed by Coulter Counter for five grain fractions: 0~125 µm, 0~80 µm, 0~56 µm, 0~40 µm and 0~25 µm. For each fraction, five different partial distributions were given (in case of glass ? four), which are the basis for the creating the joined grain composition. To assure the adequacy of the given distribution function, it is necessary to obtain the possibly best approximation of partial distributions. In the work, both the parametric and non-parametric methods of density approximations of investigated random variable. For each grain fraction the approximations were done by three traditional distribution functions: RRB, GSA and log-norm distribution functions and three methods of non-parametric approximation: Gauss kernel, Epanechnikov kernel and orthogonal Fourier series method. As the criterion of the certain distribution function acceptance, the value of the residual deviation was assumed, given by the formulae:

Jan 1, 2014

By H. G. Hedlund

Tailings in Mining Industry can be handled in different ways. Today tailings disposal is in focus by projecting new plants as well as treating and handling of old Tailings dams.The trend today is to process minerals, detoxify the Tailings and place them as dry surface Deposition or Underground Backfill. This paper will briefly present five different Mining plants in North and Central America, where Tailings have to be dewatered for dry depositing. The Metso-Sala VPA is a filter especially developed for dewatering and drying of Mineral Concentrate and Mineral Tailings. Keywords: tailings disposal, dewatering, drying, mineral tailings

Sep 1, 2012

By I. Gombköto, G. Mucsi, S. Nagy, F. Kristály, Á. Rácz, R. Szabó, B. Egyed

"Present paper deals with the investigation of a Hungarian pumice. Grinding experiments were carried out in a stirred media mill for various residence time. Ground samples were tested with laser particle size analyzer to determine and calculate the material fineness (particle size distribution and specific surface area). The lime adsorption capacity was measured to monitor the reactivity. Alkali activated cement paste was prepared using ground pumice with various alkali activators. Cube shaped specimens were produced from the paste and the uniaxial compressive strength was measured at the age of 7 days. The resulted alkali activated material structure was investigated by XRD and FTIR. The results of the experiments show that pumice fineness has strong impact on its hydraulic activity, hence on the compressive strength and the structure of the resulted binder. Relationships between grinding fineness, material structure and strength of the raw and mechanical activated material as well as the alkali activated cement were revealed. INTRODUCTION Pumice is a natural pozzolan which can be used as raw material for cementitious materials in limited portion owing to its low reactivity. Furthermore, due to its mineralogical composition it can be used for developing a new environmental friendly binder, namely alkali activated cements (AAC) which are inorganic alumina-silicate materials with good cementitious properties (Almalkawi et al., 2017; Özodabas and Yilmaz, 2013; Najimi et al. 2018; Yadollahi et al., 2015). Despite being a rather recent research topic, many papers and books have been published focusing on various aspects of alkali activation. There is great variety of raw materials that can be utilized for AAC production. However, primary raw materials are very expensive and secondary materials are heterogenous in composition. Therefore, a more homogeneous and low price silicate would help to balance the properties and cost of resulting cement and easing the way for commercialization of AAC. However, the control of its reactivity is of great importance which can be realized using mechanical activation by grinding (Juhász and Opoczky, 1990; Djobo et al., 2016; Mádai et al., 2015). The goal of present paper is to investigate the effect of mechanical activation of pumice on the properties of the resulted alkali activated cement."

Jan 1, 2018

By F. L. Wang

The purification process of water-mining carnallite and rough potassium chloride concentrate from Qinghai Salt Lake Industry Group Co. Ltd. are investigated by the preprocessing-reverse flotation technology respectively. The results show that the impurities such as calcium sulfate and sodium chloride can be effectively removed from water-mining carnallite ore and rough potassium chloride concentrate by the new technology, the low-sodium carnallite is obtained and the final potassium chloride product greatly upgraded. The content and the recovery of KCl in the low-sodium carnallite are 25.78% and 97.50%, meanwhile the content of NaCl, CaSO4 in the low-sodium carnallite are 0.56% and 0.24%, and the removal of sodium chloride and calcium sulfate in the low-sodium carnallite are 97.48%, 85.10% respectively. The KCl content and recovery of high quality potassium chloride product are 98.25%, 96.70%, and CaSO4 content is 0.40% with 83.24% calcium sulfate removed in the high quality potassium chloride product.

Jan 1, 2014

By Caibin Wu, Ali Lei, Ningning Liao, Jingsheng Ye

"The particle size distribution characteristics of ground product can influence the quality of the classification overflow and the next beneficiation indexes in an ore concentrator. Usually, the size distribution of ground product is considered to be determined by that of new ball mill feed. However, based on large amounts of data from industrial grinding circuit surveys, this paper has clearly shown that the size distribution of ground product should be determined by the size characteristics from classification underflow unless the circulating load was small, because the new feed is an important component and occupy a dominant position of mixed ore from belt feed and classification underflow. On the other hand, the correlation coefficient of the size distribution between ground product and classification underflow can be more than 0.99 when the relevance principle and its method are used. Meanwhile, their P80 value also keeps a synchronous correlation. The greater the capacity of classification underflow is, the higher the correlation of size distribution between the ground product and the mixed feed is. Furthermore, the correlation coefficient of the size distribution between ground product and classification overflow which is from primary or secondary can be 0.96. It is another discovery that the correlation coefficient is related to the type of classifier. The coefficient of size distribution of overflow and underflow from hydrocyclone is superior to that of spiral classifier, which can explain the reason why the hydrocyclone is widely used to replace the spiral classifier in concentrator.INTRODUCTION It is well knows that ball mill circuits closed with cyclones or spiral classifier such as a primary grinding circuit, a secondary grinding circuit and SAB mill have become an industry standard. Which grinding process employed will be determined by the P80 value of ground products. When the primary closed circuit ball mill or SAB that it is different throughout between the two processes was employed, a grinding fineness with -200mesh under 68% will be available. However, the secondary grinding circuit with the fineness of -200mesh above 80% can be conducted. Of course, the grinding fineness is determined by the grain size characteristics of the mineral."

Jan 1, 2018

By Graeme Heyes, Warren J. Bruckard, Michael A. Somerville

"A limited series of laboratory batch flotation tests was conducted on a suite of calcium-ferrite-based slags made in laboratory smelting trials. The smelting tests were conducted as part of a wider CSIRO Minerals’ research project aimed at developing a single-stage continuous copper-making process. The objective of the flotation work was to maximise copper recovery and iron rejection from slags made in the process.The slags treated included a reduced low copper slag, an oxidised high copper slag composite, and a very reduced low copper self-pulverising slag. The copper content of the slags varied from 2.9% to 14% Cu. The copper was present as metallic copper, oxide copper and copper-calcium ferrite phases with the phase mix related to smelting conditions. The slags were crushed, ground, wet-screened at 355 or 210 µm to remove coarse metallic particles of copper, and floated at natural pH (about pH 11) using reagents and conditions appropriate for the selective recovery of the copper phases.In rougher/scavenger tests on the reduced low copper and self-pulverising slags copper recoveries above 92% were obtained with copper enrichment ratios varying between 2.3 and 2.9. For the oxidised high copper slag copper recoveries were poorer - only 47% of the copper was recovered with an enrichment ratio of 3.3. The presence of the slow floating copper-ferrite phases in the slags limited further (selective) copper recovery. Ageing of the high copper slag, that is, a significant delay between grinding and flotation, resulted in a deterioration of the copper flotation response.The critical parameter affecting the flotation of copper from direct converting slags was found to be the form of the copper in the slag. Slags containing a higher proportion of the more desirable form of copper for flotation will provide the best opportunity for high copper recoveries to be consistently obtained, allowing improvements in the economics of the overall copper-making process."

Jan 1, 2003

By A. A. Rogozhin, E. G. Likhnikevich, S. I. Anufrieva, A. V. Kurkov

"Russia is among the nations holding the world's largest lithium ore reserves. Lithium is of key importance for the global economy. Therefore, it is important to review the current state of technology research aimed at finding the most effective solutions for the recovery of lithium from promising reserves in Russia. In Russia, a number of process solutions for the flotation of spodumene into salable concentrate (5% Li2O) have been implemented at the earlier existing complex low-grade ore concentration project. By applying these solutions, a sample of high-grade spodumene ore (4.7% Li2O) of the Greenbushes deposit was processed into a spodumene concentrate (7.5% Li2O) at a high recovery (~93%) of lithium. In Australia and Canada, new processes have been developed for the recovery of high-grade spodumene concentrates from high-grade ores using dense medium separation and flotation. For the recovery of mica lithium concentrates from low-grade ore, a flotation process was developed in Russia using the cationic collector ANP-2 and Oxifos-B as an additional collector. Low-grade ore containing 0.14% of lithium oxide was processed into a 2.7% Li2O concentrate. New hydrometallurgical lithium concentrate and ore processes are developed to avoid high-temperature (>1000 °C) roasting. The Sileach process (Australia) includes leaching of lithium ore with sulfuric acid in the presence of a halogen-containing component at a temperature of 95-115 °C. In the Russian process, lithium is recovered into an alkaline solution from low-grade spodumene ore in a pressure alkaline leaching environment (250 °C, NaOH concentration > 300 g/l). The developed concentration and hydrometallurgical methods will allow to design on their basis efficient processes for recovering salable products for hi-tech industries from lithium ores.INTRODUCTION Lithium has many unique and important properties. It is the least dense hard metal. Its alloys have some of the highest strength-to-weight ratios. Lithium has the lowest melting point of all known solid elements and is used as a highly-effective heat transfer medium. The high demand for lithium is due primarily to its unique electrochemical potential and low thermal expansion coefficient. It is used in the production of glass and ceramics, in petrochemicals, metals, medical and other industries. Since the late 20th century, the electronics industry has been the largest consumer of lithium. In recent years, lithium has attracted much attention also as an important component of ultra-lightweight electronic devices. It is also the preferred metal for electric batteries used in mobile electronic devices, vehicles, household power generators, making it an important resource for the development of environmentally friendly technology. Leading automotive brands have ambitious plans for the production of electric cars — Tesla Motors, VW, Toyota, GM, BYD Auto. It is expected that the production of electric vehicles alone will triple the demand for lithium by 2025. A sharp increase in the production of lithium is expected. Lithium has all the prerequisites to become a strategic resource of the 21st century. (Deutsche Bank, 2016; Bradley et al, 2017)"

Jan 1, 2018