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By Prasad Kopparthi, Narasimha Mangadoddy, Balraju Vadlakonda, A. K. Mukhurjee

"Column flotation (CF) is widely using equipment in mineral processing since many years. ERT technique facilitates non-invasive and nonintrusive visualization of different parameters in a column flotation. The hydrodynamic investigation of two-phase and three-phase are studied using electrical resistance tomography (ERT) coupled with pressure transducers in laboratory column flotation. Bubble size distribution, bubble rise velocity and gas and solids hold-up hydrodynamic parameters are extensively used for column flotation design and performance evaluation. The effect of superficial gas velocity, feed flow rate and slurry height in the column on mean gas holdup and its radial distribution has been analyzed for both two-phase and three phase flows. Influence of solid particles on gas hold-up and its complete elucidation of dispersion in the column are assessed. Experimental results show ERT coupled with PT is suitable as an online monitoring tool to provide useful information on the hydrodynamic parameters of the column flotation. The tomography images, which were generated using a modified sensitivity back projection algorithm, were employed to explore the parameter study in two phase flow in column flotation.Furthermore, the results also indicate that ERT is a very powerful tool for diagnosing the `inside' flow behaviour of gas-liquid two-phase and gas–liquid–solid three-phase flow in columns operating for a wide range of flow regimes.INTRODUCTIONColumn flotation is being widely used in mineral processing industries in last few decades. There has been a lot of attention in understanding the various mechanisms that take place during operation. Due to many reasons column flotation has been found to yield better performance than conventional flotation cells, particularly with fine particles (Honaker and Mohanty, 1996). Due to complex flow behavior in nature, the performance of column flotation influenced by on flow characteristics of the process, (Jena et al., 2008). Previously basic research is carried out to evaluate the effect of the hydrodynamic parameters on the performance of column flotation (Finch and Dobby, 1991)."

Jan 1, 2018

By Arman Ehsani, Ersin Yazici, Haci Deveci

The downstream treatment of chloride leach solutions of WPCB for recovery of base and precious metals by a combination of adsorption and precipitation processes were carried out. A synthetic leach solution (1 M HCl, 116.5 g/L Cl-, 30 g/L Cu2+, 2.5 g/L Fe3+, 5 mg/L Au3+, 10 mg/L Ag+ and 20 mg/L Pd2+), which was prepared based on the composition of the pregnant leach solution obtained in the previous study was used in the tests. The leach solution was initially subjected to activated carbon (AC) adsorption (20 g/L AC, 80°C) in which gold was selectively recovered over silver and palladium i.e., 88% Au recovery at 0.5 h (92% at 3.5 h) compared with only 3-6% for Ag/Pd. Copper and iron were also observed to adsorb on AC as indicated by 1.4% and 17% decline, in their respective concentrations. Following the adsorption and filtration stage, metallic copper was added into the gold-barren solution to allow reduction of Cu2+ to Cu+, and to cement Pd on metallic copper. Over the mixing period (1 h), 93% of the Pd was recovered while only 1.4% of silver was cemented onto metallic copper. Subsequently, the leach solution was gradually neutralised by NaOH to pH 4, 7.5 and 9.5 for the precipitation of copper as cuprite (Cu2O). It was found that 41% and 46% of Cu and Ag, respectively, were precipitated at pH 4 at which iron was almost totally (99.8%) removed from solution. Precipitation of copper and silver was increased further at high pHs. At pH 9.5, cumulative recoveries for copper and silver were 91% and 85%, respectively. XRD analysis of the precipitates confirmed the presence of cupric copper oxides/hydroxides and revealed that copper losses occurred at pH 4 was due to the undesired presence of cupric copper in the solution.

Jan 1, 2014

By W. Fander, Zhonghan Zhang, T. Choo, J. Enraght-Moony

"This paper describes the mineralogical and metallurgical procedures used to determine the recovery and concentrate grade of heavy oxide minerals occurring in lode deposits. The data produced also provide vital information to the processing engineers, which allows them to develop the most suitable flow sheet for that particular heavy mineral, here called an Economic Oxide Mineral (EOM). It demonstrates how results are interpreted to ensure that the proper process is used at each treatment stage to optimise recovery and to produce the most suitable concentrate at that point. Reference is made to some of the current equipment and techniques that can be used to process EOMs. The paper shows that when assessing the economic potential of a new ore body, or existing ore treatment plant, ore characterisation (OC) is necessary. This is especially true for EOMs such as cassiterite (SnO2), wolframite ((Fe, Mn, WO4,) pyrochlore ((Na,Ca)2Nb2O6(OH,F)), and tantalite (MnTa2O6) where often what seem to be small changes in ore types can have major effects on recovery and concentrate grade. Due to advances in flotation technology scheelite (CaWO4) in some deposits can be less vulnerable to OC variability than other EOM’s. Thorough mineralogical examination is shown to be the fundamental basis of correctly undertaking OC. This is because mineralogical analysis identifies each of the steps required to ensure that all ore types and zones are recognised, and differences explained, and shows the benefit in using reference samples to compare differences in the ores. The OC methods used for EOMs containing metals such as tantalum, tin, tungsten and niobium are similar; if two or more of these EOMs occur in one deposit as separate minerals, then OC will show when and how these minerals should be separated. The OC techniques described for EOMs can also be tailored for other heavy minerals, precious metals and graphite."

Jan 1, 2014

By Luís Marcelo Tavares, Juliana Segura

Comminution is one of the most energy consuming stages in a mineral processing plant. This consumption is not only related to the power supplied to equipment, but also to the indirect or embodied energy, mainly associated to materials that are worn during operation, including grinding media and liners. Until recently, preparation of iron ores in Brazil did not use comminution intensively, given that plants mainly produced sinter feed, natural pellet fines, besides lump ore. However, with decreasing ore grades, size reduction now occupies a centre stage in nearly all new plants. Technologies used up to date for processing low-grade iron ores in Brazil, in particular itabirites, have consisted of three or four stages of crushing followed by ball milling. A number of new plants are either being built or are currently being designed that will process the same type of ore, so that engineers are investigating alternative comminution routes. The paper analyses comparatively four comminution routes for processing a particular itabirite ore, many of which using technologies and layouts that have not yet been used in such application. Using data from bench scale and/or pilot-plant tests, the different routes are compared on the basis of sustainability principles in accordance to the Brazilian energy matrix, which is quite particular. It is concluded that if the sustainability criterion is used, then the optimal route would correspond to a circuit that uses HPGR technology with truncated feed followed by ball milling. Given the particular case of itabirite iron ores and specially to the one in the present work and the source of Brazilian energy, it is evident that indirect energy expenditure in comminution plays a significant role in greenhouse gas emissions when compared to direct energy, being the reason for the great attractiveness of HPGR technology. Evidently that these will have to be balanced against CAPEX, OPEX, NPV and other qualitative criteria, which can certainly lead to a different optimal circuit.

Jan 1, 2014

By G. Belardi, N. Shehu, L. Passeri

"This paper illustrates the possibilities of applying what is essentially a physical soil-washing technique for treatment of soil contaminated by semi-volatile hydrocarbons in order to reduce the level of pollution to that permitted by Italian regulations.The ground around a factory producing lubricating-oil additives consists of silty-sandy soils which have long been contaminated by heavy hydrocarbons, especially branched alkyl benzenes and naphthalene derivatives. Pollution is particularly marked in some zones (hot-spots) where Total Petroleum Hydrocarbon (TPH) values are as high as 7000 ppm. The hydrogeological characterization showed the presence of a gravely soil in a sandy/silty matrix, with big stones making preferential paths to the water circulation. The permeability is 10-5 m/s and the water-table occurs at about 3 m below ground level.After laboratory and pilot-scale studies on representative samples of in-situ soils with TPH between 2700 and 7000 ppm a flow-sheet is proposed for reducing TPH values below the acceptable limits in Italy, namely 700 ppm for hydrocarbons with more than 12 carbon atoms and less than 250 ppm for those with fewer.On the basis of indications provided by laboratory tests a mainly physical treatment process of soil washing has been developed. This is used to clean-up soil from a site polluted by non-volatile hydrocarbons with a high TPH value. The process involves a high-density attrition stage followed by wet screening at 0.5 mm and then wet-gravity separation of the minus 0.5 mm fraction in the centrifugal field.The gravity separation products consist of a light fraction with particle size finer than 15 micron which contains most of the pollutants, and a heavy coarse fraction with TPH values conforming to the limits set by Italian regulations (less than 700 ppm hydrocarbons whose structure contains more than 12 C atoms, and less than 250 ppm of hydrocarbons with less than 12 C atoms)."

Jan 1, 2003

By Hong Nguen Vu, Pavla Ovcarí, Peter Fecko, Barbara Tora

The aim of this paper is to evaluate the suitability of a coal sample from the mine Jaworzno (Poland)- locality for bacterial leaching. Laboratory research used a pure bacterial culture of Thiobacillus ferrooxidans. The results showed that bacterial leaching was effective, with a one month leaching time, when 65% of total sulphur and 85% of pyritic sulphur was leached. The paper describes changes to the pyrite during and after bacterial leaching.

Jan 1, 2003

By Liu Jie

Refractory iron ore deposits in China have various genesis and complicated types, the nature of which is poor, fine-disseminated and complex. Therefore, refractory iron ore could scarcely be separated and utilized by traditional mineral processing technology at present. It is necessary to explore deep reduction technology to change the characteristics of refractory iron ore and to use effectively the iron resource. And deep reduction technology is that on the condition of higher reduction temperature than that of magnetic roasting and stronger reducing atmosphere, the iron oxides from iron ore was reduced to metallic iron particle that is easy to dissociate. And then production from deep reduction is easily separated by magnetic separation and iron powder production was obtained finally. Deep reduction technology for Lingyang iron ore from Jilin, China, as the refractory iron ore, was studied in this paper. The influences of such main factors as reduction temperature, time and coal content were investigated. The results showed that the Metallization rate of reduction product was up to more than 90% under optimum conditions, and then higher iron grade and recovery of reduction iron powder were obtained from subsequent magnetic separation. Furthermore, the deep reduction mechanism in the system was studied by the means of XRD and SEM analysis for reduction production and Physical Chemistry of Metallurgy. Analytical results indicated that during deep reduction, the size of iron particle from deep reduction materials was increasing as reduction temperature rising and reduction time extending. The growth of metallic iron particle from deep reduction production was divided into such three stages as reduction nucleation, deep reduction and growth of metallic iron particle.

Sep 1, 2012

By P. M. Solozhenkin

On the territory Sakha (Yakutia) are located largest in Russian Federation reserves of antimony ores, Sarylakhsk and Sentochansk deposits, amaunt 10% of world stocks of antimony. A processing gold antimony ores Sarylakhsk deposit is carried out on gravitation-flotation circuit to receive gold gravitation concentrate and gold-antimony flotation concentrate with the contents of antimony over 60%. The gold-bearing concentrate is treated on place hydrometallurgicaly to a high-standard product, accepted by gold bank cash department. At the ore dressing plant reached the highest recovery in world in a concentrate of antimony 90%~91.6%, and gold 85.7%~86.0% as a result of scientific research works. Studied the mechanism of interaction cations of metals Pb2+, Cu2+, Cd2+, Hg2+, Ag+ with a surface antimonite presented on the basis of modern physical-chemical methods: nuclear quadrupol resonance, x-ray electron spectroscopy, electron spin resonance. The industrial tests of sodium dimethyldithiocarbate were carried out which confirmed high recovery of Sb and As. The leaching Sb2S3 by bacteria Desulfovibro desulfuricans and transformation antimonite in antimony trioxide by Tiobacillus Ferrooxidans improved following cyanidation . The extraction of antimony in a solution 96.0%~98.0% under optimal conditions. The common total extraction in the appropriate commodity products for gold reached 98%, for Sb 96.5%. At a combination of bioprocessing and cyanidation 85.5% of extraction Au was achieved. The new solvents for antimonite were found. Now by the Close Joint-Stock Co ?Sarylach-Antimony? industrial plant is constructed for processing Au-Sb concentrates to receive antimony trioxide and gold from various antimony deposits of Sakha (Yakutia).

Jan 1, 2014

By Jerzy A. Mielczarski

The separation of albites from quartz by flotation applying the conventional HF-methods is no longer acceptable from an environmental and health point of view. Therefore, the search for a new reagent scheme without using HF depressor for separation of albite from quartz is under consideration. The flotation behaviour of ore mineral components depends on the produced hydrophobic or hydrophilic surface coverage (chemical functionalization). Currently we have excellent possibility of monitoring surface composition and structure at molecular and atomic levels at surfaces of natural minerals contacted with aqueous solutions of collectors, activators, depressants which is vitally important for the understanding of the surface chemistry that govern selective and efficient flotation. Detailed surface characterization of minerals before and after their exposition to different solution conditions was performed by the most modern spectroscopic techniques such as developed by us the infrared external reflection technique combined with spectral simulation and X-ray photoelectron spectroscopy (XPS). It is possible to monitor surface reactions which facilitate the separation process and those which bring disadvantage to selective flotation. At first we found that a small amount of ?mineral impurity? presented in natural ore could play important role in selective hydrophobization process. It was found that the pH of solution (parameter possible to control) and natural mineral impurities present in mineral samples (parameter related to ore origin) play a very important role in the separation process. Another important parameter is interaction between minerals, quartz and albite, present in aqueous solution. The limitation of flotation separation resulting from chemistry will be discussed. Moreover, the most promising scheme for separation of albites from quartz will be examined.

Jan 1, 2014

By Z. F. Dai

The Thompson Nickel Belt, located in Northern Manitoba, Canada, contains more than 400 million tonnes of ultramafic low-grade nickel mineralization in 9 deposits. One of these, the Pipe deposit, was mined between 1971 and 1983 producing a Ni concentrate of 9% to 11% grade at 73% recovery. The operation was terminated because the poor recovery, low Ni grade and high MgO con-tent of the concentrate made processing this ore uneconomical at the metal prices prevailing at that time. Vale Inco is now re-evaluating these deposits with the intent of providing a new source of concentrate for its Thompson integrated processing facility. It is known that serpentine slimes coating on the sulphide surfaces are responsible for the poor metallurgy. Various desliming methods were tested including settling, elutriating, tabling and hydrocycloning. Hydrocycloning has proven to be the most effective means of desliming. With a feed assaying 0.70% Ni and 33% MgO, a rougher flotation concentrate assaying 22.7% Ni and 4.8% MgO was produced at a Ni recovery of 62.5% (relative to the feed to grinding), with no utilization of slimes dispersants. Another 16.6% Ni recovery was achieved in scavenger flotation resulting in a total Ni recovery of 79.1%. Batch flotation tests using various flowsheet configurations to optimize the grinding, desliming, reagentizing and flotation conditions are in progress.

Jan 1, 2014

By Luis Bergh, Francisco Díaz, Isabel Panire, Juan Yianatos, Luis Vinnett

"In this paper, industrial data of residence time distribution (RTD) in mechanical flotation cells is presented, which were measured in several plants using the radioactive tracer technique. Results include mechanical flotation cells of 100 - 300 m3, which were operated with effective residence times from 2 to 7 min. Tests were developed in forced air and self-aerated cells for liquid and solids (non-floatable) per size classes. Different RTD model structures including perfect mixer, Large and Small Tanks in Series (LSTS) and by-pass flow plus perfect mixer were evaluated and compared for single cells. The actual mixing regime was related to the effective residence times and cell designs based on the model fitting. By-pass flow percentages were typically lower than 10%, therefore the perfect mixer model plus a dead time was a suitable model structure in most cases. On the other hand, the arrangement of cells in series was effective in reducing the bypass flow observed in single cells (by-pass flow percentage greater than 5%). In addition, the N perfectly mixed reactor in series model was used to represent the mixing regime along flotation banks, which were operated with mean residence times in the range of 18 - 53 min. Results showed that N tanks in series plus dead time were closer to the actual values in most cases."

Jan 1, 2014

By Subrata Roy

The slag produced in metallurgical smelting processes contains considerable amount of valuable metals. The present work focuses on recovery of copper value by froth flotation from the copper smelter slag. The slag samples are subjected to detailed characterization including automated quantitative mineralogy using Mineral Liberation Analyzer (MLA). The MLA results indicate that a part of copper is present in sulphide form in fayalite (iron-silicate) matrix and very poorly liberated. SEM-EDX analyses also reveal that a certain part of the copper is present within the fayalite (Fe-Ca-Si) matrix, perhaps either as lattice impurity or very fine grain. The copper smelter slag is ground in wet mode in a ball mill, to particle size distribution suitable for flotation. Subsequently, the slurry undergoes froth-flotation in lab-scale Denver-type flotation cell. Various combinations of sodium iso-propyl xanthate (SIPX) and alkyl hydroxamate collectors are used along with pine oil as frother. A multiple collector dosage is required for effective flotation of non-sulphide copper phases associated fayalite and alkyl hydroxamate is identified as one of the effective collector for non-sulphide minerals. The copper recovery can be enhanced by 7-10% with a combination of collectors as compared to conventional sulphide flotation using only xanthate type collectors. However, the grade decreases to some extent. Keywords: flotation, copper slag, hydroxamate collector

Sep 1, 2012

By P. Prusti

An air dense medium fluidised bed separator (ADMFBS) is the efficient process for dry beneficiation of coal. Magnetite powder is used as medium to make required pseudo-fluid density in the fluidised bed. The dynamics behavior of coal particle with reference to its position changes from ideal solution to pseudo-fluid condition in different bed densities depending on its shape, size and density. In this study, three set of projected areas of coal particles were taken. Each set contains four different shapes i.e., cubical, rectangular prism, spherical and triangular prism with different density starting from 1.3 to 1.8 gm/cc. It is observed that coal particles having density less than the bed density, floats in different positions in the dense medium fluidised bed. Higher projected area particles float in deeper height in comparison with smaller projected area. The difference in position of coal particles in the bed increases (nearly 1 to 4 cm) as the coal density approaches towards bed density irrespective of the different shapes. Different shapes of coal particles behave in different ways based on the operating parameters like bed density, coal density and projected area of coal particle. Empirical model equations are proposed to predict the position of coal particle of each shape in the air dense medium fluidised bed system. Keywords: ADMFBS, particle dynamics, projected area, shape

Sep 1, 2012

By G. P. Wong

"In the Inco matte separation process, a slow-cooled converter matte is separated into three main discrete mineral phases: chalcocite (Cc), heazlewoodite (Hz) and a Cu-Ni-Fe alloy. After magnetic removal of the alloy, Cc and Hz are separated by flotation at pH>12 with diphenyl guanidine (DPG) as both collector and frother. High selectivity is achieved in rougher flotation. However, in later flotation stages such as in the middlings circuit, as oxidation takes place, Hz becomes highly floatable, resulting in a very poor separation between Cc and Hz.Starch and dextrin as Hz depressants were evaluated on the middlings circuit feed and showed no positive impact on Cc/Hz separation selectivity. Ethylene diamine tetra-acetic acid (EDTA) extraction showed the existence of abundant surface Cu and Ni ions. The use of triethylene triamine (TETA) or EDTA in flotation did not improve Cc/Hz separation selectivity. Microflotation tests were carried out on pure synthetic Cc and Hz minerals. It was found that Hz did not float in the presence of DPG. However, Hz conditioned with either copper sulfate solution or in contact with Cc floated readily in the presence of DPG. It is concluded that the major reason for Hz flotation in the middlings circuit is due to Cc oxidation releasing Cu ions that adsorb on the Hz surface making it reactive with DPG thus rendering the Hz floatable."

Jan 1, 2003

By F. Göktepe

Flotation gases and conditioning of the pulp with them are one of the important parameters of flotation and should be taken into account especially in the flotation of sulphide ores. In this study, flotation of pure chalcopyrite and pyrite samples were investigated with air, oxygen and nitrogen as flotation gas and the effect of conditioning with these three gases also were performed. Flotation gas was introduced for 5min, 10min, 15min and 20 min. For chalcopyrite flotation there was no drastic change in recovery when air, oxygen and nitrogen were used as flotation gas but conditioning the pulp with oxygen for 10 minutes increased the recovery about 8% where pulp potential has been approximately 25 mV with chalcopyrite and 150 mV with platinum electrodes. Conditioning with all three gases decreased the pyrite recovery by about 20% with 20 minutes conditioning. Pulp potential decreased in the presence of nitrogen while conditioning with air and oxygen increased the pulp potential for both electrodes. Generally potential increased for both minerals with oxygenation and aeration time but it was stable or slightly decreased with nitrogenation. Comparison of the gases showed that the different flotation gases do not effect flotation recovery but conditioning the pulp with these gases at pre-flotation effects chalcopyrite and pyrite flotation recovery, most dramatically pyrite.

Jan 1, 2014

By Venugopal Tammishetti

One of the important technological challenges facing the iron ore industry in India is to find commercially viable technologies to process alumina rich iron ore fines and slimes. The key to solving this problem lies in developing selective reagents for the separation of iron oxide minerals from associated gangue minerals like quartz, gibbsite, kaolinite and montmorillonite. We present in this paper our successful results on the dispersion-flocculation of three different natural iron ore slime samples from mines in India using the reagent combination of PVP dispersant and corn starch flocculent. All the three samples contained hematite, goethite, kaolinite and gibbsite minerals as confirmed by X-Ray Diffraction (XRD) analysis. The effect of various process parameters such as pH, dosage of reagents, settling time and pulp density was studied. We were able to produce a concentrate assaying 66.2% Fe and <2% Al with a yield of 67.4% by weight (with a recovery of 77% of iron values) using PVP (1 ppm dosage) and corn starch (20 ppm) for one of the slime sample (Sample I) assaying 58.2% iron (Fe) and 3.8% aluminium (Al). The XRD analysis of settled and dispersed fractions revealed that alumina-containing minerals like clay and gibbsite mostly remain in the dispersed phase while iron-bearing minerals tend to concentrate in the flocculated fraction. Keywords: iron ore, slime, flocculation, dispersion, beneficiation, starch, polyvinylpyrrolidone

Sep 1, 2012

By D. B. Sundar Ramam

In the recent past, a paradigm shift in terms of resource utilization and beneficiation practices is experienced by the Indian mining and mineral sector. Depleting iron ore reserves coupled with increasing demand for low-alumina iron ore fines to improve blast furnace performance in terms of productivity and reduced slag rate necessitate intensive beneficiation of iron ore. The task is even difficult for Indian iron ores, which are generally high in iron content and at the same time characterized by aluminous gangue. In 2006, M/s Tata Steel commissioned a state-of-the-art Jigging Plant (Batac Jig) at Noamundi Iron Mine (NIM), mainly to reduce alumina in iron ore fines. The jigging plant is incorporated in the existing beneficiation circuit to treat ?10+0.5mm size fraction of classifier fines. Over the course of plant stabilization while the plant could deliver desired ore upgradation and average product alumina, variability in terms of alumina content has been a concern. This paper discusses a case study of process parameters optimization in jigging operation to achieve consistent product quality while improving the effectiveness of gravity concentration for the beneficiation of Noamundi iron ore also to take care of inferior plant feed (input) grade year on year. The operating variables used to optimize jigging operations include bed height, pulsion and suction time, particle size, strokes, and plant throughput rate.

Sep 1, 2012

Electrochemical behaviors of synthetic and natural enargite in buffer solutions have been investigated using open circuit potential, stationary polarization and cyclic voltammetry as functions of solution pH, scan direction, scan range, scan rate and dissolved O2. The results indicate that the electrochemical behaviors of synthetic and natural enargite are very similar; solution pH, scan direction, scan range and scan rate have remarkable effects on the electrochemical behavior of enargite while dissolved O2 has little effect on the electrochemical behavior of enargite. The potential range in which enargite possesses the ability of passivation is from ?0.9 V to + 0.83 V vs. SCE. Enargite is most stable in neutral and alkaline solution around potential ?0.2 V vs. SCE. The main oxidation products on enargite surface in alkaline solution are Cu(OH)2 while the potential imposed on it is within the range of ?0.2 V to 0.33 V vs. SCE and Cu(OH)2 + S0while the potential applied is higher than 0.52 V vs. SCE.

Jan 1, 2014

By L. Bbosa

Positron Emission Particle Tracking (PEPT) and the Discrete Element Method (DEM) are used to investigate Power Draw in a laboratory scale tumbling mill. The mill is run dry in batch mode at different speeds using mono-size charge of glass beads, and is mounted along with a torque transducer to measure power. Particle tracking information from PEPT is used to reconstruct the motion of glass beads and infer the overall charge behavior, while DEM is employed to simulate particle motion and interaction, with collision mechanics calculated using the Hertz-Mindlin contact model. For PEPT and DEM data, the product of torque and average angular velocities in discrete cells are accumulated to obtain mill power. This method is found to be within statistical agreement with measured power for all tests. Spatial distributions plotted from this approach highlight the regions of the mill that draw the greatest power. A second technique to obtain Power Draw from DEM data which sums particle forces against the mill geometry is similarly found to accurately predict mill power with lower statistical error. Keywords: Power Draw, PEPT, DEM

Sep 1, 2012

By Leming Ou, Hongqiang Li, Li Zeng, Qiming Feng, Yiping Lu, Guofan Zhang

"The entrainment of sericite, especially in the presence of hydrophobic microcrystalline graphite, leads to the poor flotation selectivity of microcrystalline graphite ore. In this paper, it was attempted to reduce entrainment of sericite, selective flocculation of microcrystalline graphite using polyethylene oxide (PEO) was studied. For the sake of flocculate microcrystalline graphiteselectively, flocculation tests of microcrystalline graphite and sericite single mineral were conducted.To investigate the effect of selectiveflocculation of graphite on entrainment of sericite in microcrystalline graphite flotation, batch flotation tests of graphite single mineral and artificial mixtures were conducted; theflocculationselectivity of PEO of different molecular weight (M.W.) was alsocompared inartificial mixture flotation. Inthe end, the structure of flocs in different particle size was studied by sieve analysis and microscopic observation. The resultsindicated that PEO of molecular weightone million andtwo million could selectivelyflocculate the microcrystalline graphite and reduce the entrainment of sericite at a lower dosage; further increasingof the dosage of PEO(M.W. ?two million) led to hetero-flocculation between graphite and sericiteandmadethe ash content of concentrate increase again."

Jan 1, 2014