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By Anand Anupam

Conventionally, beneficiation in Indian coal washeries is done using dense medium cyclones upto the size of 0.5 mm. Coal fines below 0.5 mm size are extensively cleaned by froth flotation. There are many problems associated with flotation of Indian coals, e.g. poor response of coal particles to flotation, misplacement of fine high-ash material to froth, loss of good quality coarser coals (within ?0.5 mm size range) to tailing streams, high cost of reagents, water contamination, etc. In addition, sophisticated control systems for efficient management of the fines washing circuits are needed. Thus, there is an escalation in the capital and processing costs. In view of these facts, studies were carried out to establish the cleaning potential of dense medium cyclones for beneficiation of fine coal down to a size of 0.1 mm, essentially aimed at reducing the load on fine coal circuits mainly comprising of flotation. ROM Coal (in the size range of -2 + 0.1 mm) obtained from JogtaOCM,Jharia coalfield possessing relatively low ash (about 21%) was used in this study. Purpose for using low ash coal was to further lower, its ash content down to 8% which is comparable, to the ash content of imported coal. After characterization studies, nine tests were carried out on the prepared feed sample for its beneficiation using a dense medium cyclone of non-conventional design (very low cone angle and longer barrel), by varying its apex and vortex finder diameters. The efficiency of separation has been evaluated using Tromp curves. It could be established that it is possible to produce a clean coal of 8% ash, with Ep value of 0.044, using Jogta OCM coal by treating it in a modified dense medium cyclone. These results are summarized and discussed in detail in the present paper. Keywords: dense medium, cyclone, fine coal, beneficiation, Jharia coal

Sep 1, 2012

By D. Izerdem, S. L. Ergun

"Beneficiation of the finely disseminated valuable minerals has been an important issue in the mineral processing field for decades. Treatment of the fine materials becomes difficult with using traditional gravity concentration equipment. To investigate particles behavior in a commonly used conventional gravity concentration equipment, a comprehensive study was undertaken in a plant, which only consists spiral concentrators as concentration units. Spiral concentrators require relatively coarse material to effectively concentrate valuables as well as offering greater simplicity for especially higher grade feeds. In the current work, chromite ore (d80 = 290 µm) was treated using spiral concentrators and the performance of spiral concentrators was examined. The design of the separation circuit comprised of rougher, cleaner and scavenger stages. The resultant fractions produced by each separation unit and feed samples were analyzed to identify the particle sizes, mineral liberations and chromite content. Grade and recovery comparisons were presented in size-by-size manner and the separation tendency of the particles were determined in varying feed rate conditions. The overall results indicated valuable mineral loss of the plant and alternative suggestions were pointed to reach the target values.INTRODUCTION Spiral concentrators are simple low energy consuming devices, which are generally used for pre-concentration of the minerals. They are cost efficient equipment separating minerals mainly based on density differences of the minerals. The adjustable splitters remove the concentrate as the remaining stream continues to move down the spiral. Separation on a spiral is achieved through a combination of forces that act on particles as they move down the trough of the spiral. The main forces known to act on the particle on a spiral are the gravitational force, centrifugal force, hydrodynamic drag and Bagnold force (Atasoy and Spottiswood, 1995; Burt, 1984). Since the classification in a spiral concentrator depends on particle specific gravity and size, the operation of spirals can be analyzed using size recovery or partition curves (Bazin et al., 2014). Spiral concentrators possess multiple design parameters for a variety of applications. The design parameters that govern the performance of the spiral can be identified as the diameter, number of turns, slope, pitch and the profile of the trough. For an efficient separation on a spiral concentrator, the properties of the slurry flowing on a spiral, ore texture, percent solids and flow rate play important roles."

Jan 1, 2018

By E. Scheepers

The process under investigation is the large-scale production of phosphorus. Gravity delivers the feed, consisting of pelletised apatite, coke and silica (in the form of gravel), to a submerged arc furnace through ten, evenly distributed feed chutes ensuring constant packed bed volume. Within this packed bed, solid-gas reactions are dominant in generating the gaseous product. These reactions, as well as the packed bed itself, were the main focus in the development of a fully three-dimensional CFD model in which fluid flow and energy distribution are represented as functions of chemical reactions, heating and melting of material and heat losses. The associated complex process interactions necessitated the creation of two process-specific, user-developed models that integrates accurate, thermodynamical data with computational fluid dynamics calculations, as well as reaction kinetics. The boundary conditions are provided by industrial measurements, and the key model parameters of reaction kinetics were obtained by laboratory experiments for different feed mixtures. Thirteen temperature probes installed inside the refractory lining of the furnace especially for this project allowed parameter fitting and validation to be guided within the constraints of real measurements. Simulation results reveal a thin, dome-shaped, reductive zone where conditions are optimum for maximum phosphorus recovery, as well as evidence of a thermal reserve zone similar to an ironmaking blast furnace located above this reduction zone. When compared to actual operating data, the CFD model can successfully predict the specific power consumption, the phosphorus in the slag and gaseous product flowrate under given operating conditions. This CFD model forms part of a larger project that aims to create a Dynamic-CFD hybrid control model.

Jan 1, 2014

By B. Rezai

Impeller speed and air flow rate are two influence parameters on coarse particles flotation and collision and attachment angles. So, in this research recovery of quartz particles at four diameter classes carried out and collision and attachment angles investigated. It is found that maximum flotation response was observed at Impeller speed of 1100 rpm and 15 L/h air flow rate. For either more quiescent (Impeller speed<1100 rpm) or more turbulent conditions, flotation recovery decreased steadily. Furthermore, amount of attachment probabilities is calculated using suitable equation. According to this study, with decreasing particle size, air flow rate and impeller speed, probability of attachment increased and with increasing contact angle, probability of attachment increased. For various air flow rates, impeller speed and particle size, collision angle has been calculated. Maximum collision angle obtained at with 15 L/h air flow rate, particle size of 256 µm and impeller speed of 700 rpm and minimum collision angle obtained at with 45 L/h air flow rate, particle size of 545µm and impeller speed of 1300 rpm. In this study, maximum attachment angle, , was around 75 L/h air flow rate, particle size of 256 µm, contact angle of and impeller speed of 1300 rpm and minimum attachment angle was with 15 L/h air flow rate, particle size of 545 µm, contact angle of and impeller speed of 1300 rpm. ?82.72 ?64 cr??33.39 ?90 ?13.0 ?60

Jan 1, 2014

By V. Beskrovnaya

Oxidised lead-bearing ores of one of the deposits of Russian Federation represent complex raw materials. Its main components are lead (3.2%~4.7%), silver (100~300 g/t) and zinc (0.7%~0.8%). The sample is characterized as oxidized ore. Silicium (37.4%) and iron (33.2%) are its main com-ponents and present in oxide form. Lead presents as cerussite, anglesite and plumbojarosite have a subject role. Zinc occurs as smithsonite. Certain part of zinc is associated with manganese minerals (psylomelane-pyrolusite). Close association of valuables with manganese-iron mineralization is ob-served. Silver is present in native finely dispersed form. Three options of flowsheets have been developed to recover valuables from the ore. 1. Hydrometallurgical flowsheet of lead and silver recovery on the basis of chloride leaching. 2. Flotation. 3. Flotation -flotation tailings cyanidation. Hydrometallurgical flowsheet of ore processing on the basis of chloride leaching (option I) al-lows to achieve recovery into saleable products of 68.8% Pb and 70.2% Ag. High quality lead concentrate with 71.7% lead and 700 g/t silver was obtained during flotation (option 2). Flotation tailings cyanidation (option 3) provides 24% silver additional recovery (from the ore). Technical and economical evaluation demonstrated that ore flotation (option 2) and flotation with subsequent flotation tailings cyanidation (option 3) are the most cost-efficient. Option II is the most cost-efficient variant of 1 tonne of ore processing (9.05 US dollars).

Jan 1, 2014

By A. Zangooi

Frothers play major roles in flotation through interaction at the air/water interface. A method of determining frother adsorption density (coverage), i.e., the number of molecules per unit of interfacial area of bubble surface, is introduced. The technique is based on mass balancing the frother around a continuously operated bubble column. The mass balance requires collection of bubbles before they burst. The column design and operating procedure is reviewed, and some preliminary results discussed. Keywords: flotation, colorimetric frother analysis technique, frother mass balance, coverage

Sep 1, 2012

By Pat Turner, Abrie Schutte, Tim Olson, Roman van Ommen

The key to good cyclone circuit design and operation in primary grinding circuits will be discussed. Cyclone circuit design issues include individual cyclone design features such as cone angle or materials of construction, as well as orientation and manifold configuration. Cyclone performance issues as well as maintenance considerations will be detailed in an effort to achieve consistent cyclone operation with minimal coarse solids in the overflow. The new gMAXTM. cyclones will be introduced and evaluated using operating circuit data in closed circuit grinding applications

Jan 1, 2003

By T. Fraszczak, U. A. Peuker, T. Mütze, T. Leißner, A. Wikedzi

The present paper highlights on the comminution and liberation properties of two greisen-type lithium bearing-mica ore blends (L1, L2) subjected to different breakage devices; cone crusher (CC), roller crusher (RC), rotor beater mill (RBM) and a screen mill (SM). The particle size distribution (PSD) of the products from each device were evaluated to search for an appropriate PSD model using Gates-Gaudin-Schuhmann (GGS) and Rosin-Rammler (RR) functions. In order to determine an appropriate function that could better describe the PSDs from the devices used, the coefficient of determination (R2) was used as a criterion. In terms of mineral liberation studies, due to costs, only products from rotor beater mill (RBM) were examined through an automated scanning electron microscope (SEM) technique. It was found that RBM, RC and SM products were better described by the RR model than the GGS model with higher R2 values (0.97-1.0). However, cone crusher products for L1 were better described by GSS model, while that for L2 were better described by RR model. In terms of the spread of size distribution as indicated by RR model parameters, RC products were more uniformly distributed compared to products from other devices, for both ore blends. Also the RBM products were more scattered than products from other devices. The results indicate that the composition of individual ore blends affected the comminution products PSDs since different PSD model parameters were obtained for samples comminuted by same devices. The modal mineralogy of the two ore blends indicate both to be rich in quartz, topaz, zinnwaldite and muscovite. Further, the result indicate that, for both ore blends, zinnwaldite phase is more enriched in the fraction < 250 µm. Moreover, better liberation of zinnwaldite is observed for L1 compared to L2. This could be explained by differences of the two ore blends in three aspects; the nature of mineral association, reduction ratio of the fractions analysed and the nature of spread of the size distribution.

Jan 1, 2018

Over the last four years, a major collaboration between CSIRO-MIS, UCT and JKMRC has produced two significant advances in the modelling of comminution. The Universal Comminution Model (UCM) uses parametric SPH/DEM studies of comminution and separation devices to calibrate mathematical models which are suitable for industrial use. The Virtual Comminution Machine (VCM) uses the same computationally intensive techniques based on discrete element modelling (DEM) and smoothed particle hydrodynamics (SPH) to directly model any comminution device by combining a CAD file describing the machine design, sophisticated ore characterisation methods and advanced particle scale flow and breakage simulation. The UCM should be useful for process analysis and optimisation of existing equipment and circuits while the VCM should make a major contribution to the evaluation and development of new machines. Progress to date is encouraging with several processes now well described using DEM. This paper also reports on progress towards combining DEM for particles with a realistic slurry model based on SPH for different types of comminution devices. In due course, this combination should allow accurate modelling of particle transport in wet mills to be combined with DEM models.

Jan 1, 2014

By W. Y. Liu

A sulfate reducing bacteria strain (named SRB-A) is isolated from sanitary wastewater by using anaerobic technology of Hungate. The studies of physiological, biochemical and phylogenic of the isolated strain are carried out. The colonies of the strain are circle. Cells that are slightly curve, rod-shaped with the size of 1.0~3.0µm in length and 0.8~1.0µm in width. The optimum temperature for growth of the organism is 30~35°C. The optimum pH for growth is 7.0~7.2. Strain SRB-A can use sodium lactate, sodium citrate, sodium acetate, ethanol and acetone as electronic acceptor but can not use glycerol, glucose, sucrose and sodium oxatate. The isolated strain reduces sulfate, sulfide, sulfur, DSMO and FeCl3, then nitrate and nitrite can not be used. The strain has higher tolerance of NaCl, and the optimum growth occurs in the presence of 5%~8% (w/v) NaCl. The result of 16S rDNA sequence analysis indicates that the isolated strain SRB-A is the most similar to Desulfo-vibrio halophilus, and the level of the similarity was 98.4%. The isolated strain has a stronger ability to reduce sulfate, and the level of sulfate reduction reaches about 90% at pH 7 in the liquid medium. 0.6 to 0.8 is the optimum COD/SO42- in the medium for the isolated strain to show better ability to reduce sulfate. At pH 2.5 the level of removal of Cu2+ reaches 96.3%. Under the condition of pH 3.5, the level of removal of Zn2+ reaches 92.6%. When the value of pH reaches about 7, the Fe3+ is precipitated by the formation of Fe(OH)3, and Fe3+ is removed up to 89.4% at last.

Jan 1, 2014

By V. V. Ramarao

Grinding and flotation operations deliver better-optimized performance when they are monitored and controlled in terms of the minerals they process rather than as elemental assays that we usually refer to. Mesh of grind (MoG), liberation, floatability, grade and recovery etc. are more directly related with the minerals and their characteristics rather than the elemental assays. Hence, online analyzers only provide the status for the grade and recovery of metals and not the direction for the enhanced performance. Central Research and Development Laboratory (CRDL) and Rampura Agucha Mine (RAM) mill section of Hindustan Zinc Limited (HZL) jointly undertook various survey campaigns in the grinding and flotation process streams. Grinding circuit surveys involved systematic collection of representative samples from different grinding circuit streams under stabilized conditions for monitoring particle size and process mineralogy details across individual grinding circuit. Flotation survey campaigns included study of fort-nightly / monthly composites of un-ground flotation feed, lead concentrate, lead tailing, zinc concentrate and final tailing samples for size distribution details in the 150 ? 38 micron range, chemical analysis and mineral characterization through optical microscope. This work details the investigations results and their implications on beneficiation performance. Grinding circuit survey campaigns supported the task for modifying the originally installed configuration of hydrocyclones, to limit recovery losses due to excess grinding of soft galena and related silver minerals. Flotation plant surveys helped in adjusting flotation circuit process parameters to reduce loss of liberated and floatable galena and sphalerite grains into their respective tailing streams. Keywords: grinding circuits, survey campaign, hydrocyclones, mineral liberation and locking, lead-zinc ore beneficiation operation

Sep 1, 2012

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 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 H. C. E. Regidor, M. G. Mena, H. D. Mendoza

Cyanidation leaching has always been the most widely used method of extraction of gold in the world. Its low cost and high selectivity has made it the prevalent choice for mining companies, and even for the small scale miners. The use of cyanidation in small scale mining has been an issue for the longest time, mainly due to the operation’s lack of safety measures, health care or environmental protection regulations. In pursuit of more environmentally friendly methods of extraction of gold, several alternative extraction processes have been explored, one of which is chlorination. The chlorination process is advantageous when recovering refractory gold locked in sulfides, as it simultaneously oxidizes the surrounding sulfides as well as dissolve any gold exposed to the lixiviant. This research was conducted to study and develop a safer alternative way of extraction of gold from small scale mining operations, without compromising the recovery of gold. In this study, a variation of the chlorination process has been employed for a refractory flotation concentrate wherein the majority of the gold is locked in sulfides. Hypochlorite-bearing reagents were used as the leaching reagents, and the method developed was a 2-stage leaching process in differing pH conditions. The solution pH and Eh was monitored all throughout the runs, and aliquot samples were taken at time intervals for analysis. SEM-EDS imaging was used to determine the mode of oxidation of the chlorination process, and the image obtained showed that the oxidized layer followed the shape of the particle, suggesting a diffusion controlled process. XRD analysis of the sample before and halfway through leaching further confirmed the oxidation of the sulfides. The amount of gold dissolved in the aqueous solution and in the solid residue were used as the basis for the determination of the recovery of gold. An average gold recovery of up to 84.38% was obtained using the developed method in just 7 hours of leaching. This method was then employed for a pilot-plant scale test and a gold recovery of 79.2% was obtained.

Jan 1, 2018

By Lilong Huang, Xiang Zhao, Yuan Long, Li Shi, Guowang Zhang, Xiao Xiao

"Vertical stirred mill is expected to upgrade the iron concentrate by further size reduced efficiently. The tests for fine grinding of iron ores from Baiyun Boyu iron ore, Baogang Goup, iron contained tails of Cu-Pb-Zn ore in Wanguo ore, Xiaoshalong iron ore in Qinghai were conducted in this paper. The data of appropriate grinding fineness, optimum grinding conditions and energy consumptions obtained from the test were used to select appropriate scale of the vertical stirred mill, optimize the parameters of the grinding process and upgrade the iron concentrate of the iron ore dressing plant or design the new ore dressing plant. Based on the laboratory tests, industrial system of regrinding and magnetic separation with vertical stirred mill as the regrinding mill in micro-fine grained iron concentrate upgrading has worked out very well. The grade of Fe in iron concentrates has increased to about 65.0% (by 4.0~6.5 point percent) when the iron concentrates were reground to designed fineness and the specific energy is 5.0~7.1kWh/t. The system of regrinding and magnetic separation is proven to be eco-efficient in iron concentrate upgrading."

Jan 1, 2014

By S. V. Petrov

Oxidised ores possess a significant share among raw materials which are treated by traditional plant technology. As usual, the chemical composition of these ores is rather simple. They can be treated by standard cyanidation. In this case, the resulting wastes have an alkaline reaction. Their toxicity (in case of ROM ores treatment) is mainly due to the presence cyanides and thicyanades. The tails are dexotified before disposal. However, chemical reagents used cannot provide a complete removal of toxic matters. In this case, this disposal poses a real problem because these tails contain a significant amount of CN and SCN and their behavior during a long disposal should be studied. A sample with a simple composition taken from cyanidation tails was studied. The sample presents itself the quartz low-grade sulfide gold-bearing ore. Sulfide sulfur content is 0.15 per cent,and heavy metals content is 0.023 per cent. Method of testing is described further. The samples were tested under moderate and subarctic climates based on the difference in temperature, moisture and the degree of oxygen saturation of air. The pH of water extract, which was prepared using distilled water, was also determined besides measuring the concentration of cyanides and thiocyanates. This demonstrates the stability of sulfide-bearing minerals in the disposed tails. The tests time was 820 days. It was found that the alkalinity of the samples tends to lower under oxygen-containing conditions in proportion to temperature increasing. Under conditions without oxygen, the pH lowering is minimal. Mineralogical and chemical properties of the ore material give positive alkaline balance during testing. The solid sample contains enough alkaline agents to compensate dangerous to the environment reactions of sulfides oxidation.pH value maintaining under alkaline conditions witnesses to the possibility of a safe disposal of this tails type. The largest amount of cyanide is removed under acid-containing conditions at positive temperatures. The higher the temperature is, the more intensive destruction is. At the moment, of finishing the tests at 20ºC temperature, the residual concentration of CN- was 0.06-0.10 mg/L, at 4ºC, the residual concentration of CN- was 0.2-0.3 mg/L and at -18ºC was 0.4-0.6 mg/L. Under conditions with the lack of oxygen, cyanides can be removed within the first 200 days. Then, this process fully stops. A stabilized cyanides concentration is 3.0-4.0 mg/L. The presence of sulfides increases the content of cyanides at the initial stage of testing. Then, the concentration of SCN at the temperatures above 0ºC decreases rapidly under oxygen-containing conditions. At the rise of temperature, a narrowing of peak at the curve and an increase of the destruction process are observed. At the end of testing, the concentration of thiocyanates in the moisture of wastes was 0.2-0.3 mg/L at the temperature of 20ºC and 0.5-0.6 mg/L at the temperature of 4ºC, respectively. The concentration of thiocyanates contained in a frozen sample and media which have the lack of oxygen stabilizes at a certain level owing to sulfides and does not practically change further. It was found that detoxification of oxidized cyanidation tailings with a simple composition which were disposed as filtration cakes can be possible. In this case, the concentration of cyanide in the moisture of disposed wastes can have a close value or equal to actual concentration during cyanidation. In order to meet the requirements of high ecological safety standards and successful detoxification process of toxic compounds in the tails under the action of natural factors, the conditions of air aeration with oxygen for the whole volume of tails are required. Keywords: cyanides, thiocyanates, storage, detoxification, flotation cakes

Sep 1, 2012

By Chen Dong, Yang Wenwang, Liu Zhineng, Tan Ming

"To meet the requirement of production expanding of a certain potash ore dressing plant in Xinjiang province of China, industrial test of XCF/KYF-50 flotation machine was performed to process potash ore. The tank of XCF/KYF-50 flotation cell is in U shape, 50m3 of volume. This flotation cell is able to suck slurry by itself, and can be arranged on the same base level without high difference. In this paper, two sets of XCF/KYF-50 flotation cells were adopted to replace 8 sets of 8m3 flotation cell in original flow sheet of flotation. Aeration rate, slurry suction capacity, power consumption and slurry suspension are investigated. The measuring results show the slurry suction capacity of the flotation cell is larger than 12.80m3/min and solid particles disperse evenly in the tank without deposition. Industrial test result indicates that the final grade of concentrate is 26.69% at the recovery of 90.95%, which is up to expected performance. XCF/KYF-50 flotation cell has advantages of high flotation efficiency, significant energy saving and reasonable configuration of automatic control system, this kind of flotation cell can meet the requirement of production expanding of this potash ore dressing plant well."

Jan 1, 2014

By J. Jirestig

Poor grade iron ore, suffering from interlocked particles, is challenging for any process technology. Extensive grinding is required to attain liberation and to yield concentrates of desired grade. Fine particles are commonly very difficult to retrieve, resulting in value losses to tailings. In this paper, results from HGMS tests on fine to ultra fine hematite are reported. Special emphasis is put on slurry flow velocity. If hydrodynamic drag force, i.e. flow velocity is not controlled and kept constant during the feed and/or rinse phase, fine particles will not be captured and retained. Flow velocity control and constant flow during separation are achieved by discharge valves/slots and parallel magnetic flux and pulp flow configuration. In tests, increased slurry velocity from 178 to 268mm/s caused Fe-recovery to drop from 78 to 48%. Proving that flow velocity control is essential to optimise grade and recovery. HGMS tests at the Metso mineral laboratory on natural (weathered) hematite fines clearly show a large potential to recover valuables, currently lost to tailings. Even at particle size distributions as low as D80=35µm concentrate grades and recovery are very satisfying. Two samples, in-plant desliming rejects (D80=35µm and 50µm) where significantly upgraded and recovered, from 45,3% to 63,6% and 53,0% to 63,4% Fe, at 73,8% and 81,1% Fe-recovery respectively. Other types of in plant rejects, such as gravity separation or WHIMS tailings, also benefits from HGMS separation. Gravity separation reject, 55,8% Fe, at D80=500µm was tested, yielding a magnetic concentrate of 60,8% Fe at 50,7% Fe-recovery. The results indicated that the hematite was poorly liberated. HGMS separation was repeated after a regrind to D80 150µm, resulting in a magnetic concentrate of 62,8% Fe at 75,9% Fe-recovery. The tests prove that hematite, presently lost to tailings streams, can be recovered by HGMS and thereby significantly contribute to improved overall economy. Keywords: HGMS, magnetic separation, recovery of ultra-fine hematite, Indian hematite ores, hematite recovery from tailings

Sep 1, 2012

By W. Y. Gao

Diaojuntai Concentrator adopts a process of low intensity and high intensity magnetic separation and anionic reverse flotation, with all the concentrate from magnetic separation being treated by re-verse flotation to obtain a qualified final concentrate. The high reagent cost and energy consumption of the reverse flotation have led to a high concentrate cost. The test research on reducing the feed to the reverse flotation was therefore carried out. Based on the study of the properties of the ore sample, the concentrate of low intensity magnetic separation was screened at 0.076 mm and the undersize fraction was concentrated by concentrator with gravity and magnetic field. The results show that a qualified concentrate can be obtained at various rising current rate, pulsating current and power-on style, with the operation rate over 85%. The extraction of some part of qualified concentrate from the concentrate of low intensity magnetic separation by magnetic separation can reduce the feed to the reverse flotation by about 35% so as to reduce the production cost of concentrate.

Jan 1, 2014

By S. N. Titkov

Flotation of halite?new, recently begun to be developed method of processing water-soluble minerals. In the paper the results of studies of the new directions of the application of halite flotation are represented. Studies on the flotation separation of halite and sylvite in the salt solutions of different composition, showed the possibility of the halite flotation realization not only in the carnallite brine (content of salts: MgCl2 25.5% ~ 26.0 %, chlorides of potassium and sodium 5.5% ~ 6.0%), but in the sylvinite brine, which contains 10.5% ~ 10.8% KCl and 20.5% ~ 21% NaCl. On the basis obtained results are carried out experiences on the halite flotation from samples of the converted products of sylvinite ores by flotation method and by the method of the electrostatic separation. It is shown that in all cases it is possible to accomplish effective and selective recovery of halite. The halite recovery into the foam product of flotation makes 80% ~ 90% with an increase in the content of KCl in the cell product of flotation in comparison with a composition of flotation feed: - from the product after the electrostatic separation of ore ? from 76% KCl to 95% ~ 97% KCl; - from the small fraction of ore (- 0.25 + 0 mm) ? from 30% KCl to 60% ~ 64% KCl; - from middlings (intermediate products) of sylvite flotation from 16% ~ 31% KCl to 45% ~ 80% KCl. An increase in the molecular weight of alkylmorpholines increases the selectivity of the flotation separation of halite and sylvite. Halite recovery from small size fraction of ore or/and middling of sylvite flotation with the sub-sequent joint filtration halite product and tails of sylvite flotation allows to reduce considerably expenses for manufacture potash chloride from Pilot-industrial tests for the first time confirmed the possibility of the effective halite flotation in the saturated solutions of chlorides of potassium and sodium (sylvinite brine).

Jan 1, 2014