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Pyrite and Chalcocite have different crystal lattice, with different dissolution rate in the same ac-id solution, the dissolution rate of which is accelerated with increasing redox potential, they are leached simultaneously at high redox. In present work, some methods are adopted to inhibition the bioleaching of the pyrite by controlling the redox potential. Shake flask experiment shows that the dissolution of pyrite and chalcocite is slow as redox potential <700mV(Eh), but the dissolution in-creases exponentially as redox potential>800mV(Eh), they have good selectivity at redox potential between 700mV and 760mV(Eh).

Jan 1, 2014

Large quantities of electrical and electronic equipments are discarded at the end-of-life (EOL) in different industries and house hold applications. Their disposal is creating a challenging environmental problem due to the presence of hazardous metals and non-metallic materials such as epoxy resins and plastics including brominated flame retardants etc particularly in printed circuit boards (PCBs). The recovery valuable metals and separation of hazardous materials from discarded PCBs will not only conserve the resources but also mitigate the environmental pollution. From a recycling point of view for a sustainable society, the studies have been carried out for the recovery of metal values from EOL-PCBs of DVD video players (DVD-PCBs) and vacuum cleaners (VC-PCBs) using physical beneficiation techniques. The concentration of the metals was found to be 43 wt.% and 30 wt.% in DVD-PCBs and VC-PCBs, respectively. Both the PCBs samples were ground in a hammer mill after cutting and their size analysis was carried out. The liberation of the metallic and non-metallic fractions was investigated under the microscope for different size fractions. Subsequently, froth flotation and pneumatic air separation studied were carried out to enrich the metallic constituents. The studies showed the enrichment of metals in coarser particles (-1000 +150 µm) and non-metals in the finer particles -150 µm following mechanical recycling process. A grade of 90 % with 75% recovery was achieved by froth flotation, but lower grade and recovery of 75% and 65% respectively were obtained by pneumatic separation with -1500 µm powders for DVD-PCBs. A grade of 86% metal was obtained by froth flotation with 76% recovery in the case of VC-PCBs. The lower grade of 65% with almost same recovery was achieved during pneumatic separation studies due to lower concentration of metals in VC-PCBs. The studies also showed that the froth flotation is an efficient process for enrichment of metals from both the samples. The process is environmental friendly as no chemicals are employed during entire processing. The higher recovery and grade of metal is expected to be obtained on a large scale beneficiation process.

Sep 1, 2012

This investigation addressed two products from Samarco?s desliming circuit, namely underflow of cleaning cyclones and underflow of desliming cyclones. The roughing cyclones receive the product of the grinding circuit. The underflow of the roughing cyclones feeds the cleaning cyclones and the overflow feeds the desliming cyclones. The underflows of cleaning and roughing cyclones are combined to constitute the feed of the flotation circuit. These fractions were characterized and a new process route was proposed. The size distributions of the two fractions are quite different. The desliming cyclones underflow presents higher amount of material passing 37 µm and of ultrafines as compared with the cleaning cyclones underflow. Laboratory scale flotation tests were performed in a mechanical machine aiming at comparing the performances of these products in the presence of the reagents employed in the concentrator. Pi-lot scale column flotation yielded enhanced iron recoveries for both products in comparison with those achieved in laboratory mechanical cells, without impairing the concentrate quality. The in-crease in recovery was more significant in the case of the cleaning cyclones underflow. These results suggest the use of columns in the flotation of this product aiming at increasing the overall iron recovery. Nevertheless the results also showed that the use of columns to float the cleaning cyclones underflow should also be taken into consideration. Despite the fact that this product requires a longer residence time in order that the required silica content in the concentrate is achieved, the indication is that higher levels of iron recoveries might be reached. The conclusion is that, due to the different behaviors of the two products, they should be processed in distinct circuits to maximize the iron recovery.

Jan 1, 2014

By X. Xu

This paper presents the research on comprehensive recovery of copper, gold, silver and iron from tailings from high-intensity magnetic separation in Tonglushan Mine. By comparison of experimental results of three flowsheets?regrinding of tailings-conventional sulphidizing flotation, water heating enhanced sulphidizing flotation and acid leaching-sulphidizing precipitation flotation, it is concluded that regrinding of tailings-conventional sulphidizing flotation flowsheet has the best economical and technical performance and practicability. Regrinding and using the mixture of sodium hydroxamate and sodium xanthate are two of the notable technical characteristics of this best flowsheet. The system for comprehensive recovery of tailings has been put into production. It not only can recover the valuable minerals from tailings in Tonglushan Mine, but also can reduce the environmental pollution, resulting great economic and social benefit.

Jan 1, 2014

By T. Saastamoinen, N. Heino

"The Olserum deposit located in Sweden hosts an indicated mineral resource of 4.5 million tonnes over 0.6 % TREO, exhibiting a relatively high proportion of HREE. Beneficiation of a composite drill core sample was studied in the flowsheet of magnetic separation of magnetite followed by REE minerals flotation. Primary testwork was performed to select an effective REE mineral collector and then the parameters of grinding size, collector and depressant dosages, slurry pH and density, conditioning time and temperature, and flotation time were tested for optimization of REE flotation. The studies of process mineralogy showed that the REE-bearing minerals monazite, xenotime and apatite were successfully enriched from the concentrations of 0.6%, 0.31% and 2.6% in the feed to those of 17.0%, 8.9% and 65.0% in the concentrate with the recoveries of 79.0%, 81.3% and 71.0%, respectively. INTRODUCTION The Olserum deposit located in Sweden comprises lower grade placer material with higher grade apatite-magnetite veins and hosts an indicated mineral resource of 4.5 million tonnes at 0.6% TREO, with 34% of that being HREO (EURARE, 2017). The mineralization was studied by Andersson et al. (2018). At early stage, high-temperature REE is characterized by abundant monazite-(Ce) and xenotime-(Y) coexisting with fluorapatite and subordinate amounts of (Y,REE,U,Fe)-(Nb,Ta) oxides. During a subsequent stage, allanite-(Ce) and ferriallanite-(Ce) formed locally, partly resulting from the breakdown of primary monazite-(Ce). Alteration of allanite-(Ce) or ferriallanite-(Ce) to bastnäsite-(Ce) and minor synchysite-(Ce) at lower temperatures represents the latest stage of REE mineral formation. The mineralogy and beneficiation were firstly studied by GTK Mintec in the EURARE project for a composite drill core sample. The Olserum is considerably smaller than the Norra Kärr deposit. However, REE are hosted by xenotime and monazite, both of which have well established processing pathways, and meanwhile, the Olserum has a high contribution of the high value critical REE (dysprosium (Dy), yttrium (Y), neodymium (Nd) and terbium (Tb)), which provide a unique advantage for its exploration (Tasman Metals Ltd., 2013). In this paper, flotation of the REE minerals in the Olserum ore was studied."

Jan 1, 2018

"The most common comminution testing methods and the associated analytical procedures provide average values and do not capture the variability in breakage behaviour of ore particles. Therefore, plant designs and process performance predictions use only average values; hence, the inherent variability that exists within an ore domain translates into variation in process performance. This paper discusses the drawbacks of using the average approach to impact comminution testing and plant design. It considers breakage intensity from a single-particle testing perspective. To clearly delineate breakage behaviour of each broken particle, an experimental technique using the JK Drop Weight Tester (JKDWT) has been developed by which actual comminution energy per particle and breakage intensities were measured directly. Four different ore types in ranges of -45+37.5mm and -22.4+19mm were tested at three energy levels of 0.1, 0.25, 0.5 kWh/t and six energy levels of 0.5, 1.0, 2.0, 2.5, 3.0 and 4.0 kWh/t, respectively. Sieved results of individual particles, notionally, t5, t10 and t50 showed the presence of large variability in their breakage behaviour, which was not captured by the standard JK impact breakage testing approach. In addition, the original JKMRC model is modified by introducing a third parameter, ‘Emint10’ which describes the minimum energy required to produce the minimum percent t10 for a given ore type and its significance is tested statistically. These results suggest that the analysis of breakage variability may lead to more robust plant designs. Perhaps, the most useful insight from this work is to extend single particle breakage model developed by JKMRC to a comminution percentile envelope which may provide a more detailed picture of ore breakage variability. INTRODUCTION Rock consists of mineral grains, pores, voids and cracks that may differ markedly in size/density which make rock a heterogeneous material by nature. The degree of heterogeneity in rock affects its breakage characteristics where these variations increase local differential strain and stress within a stressed rock (Scholz, 1968) and it depends on many factors such as rock genesis, alterations, etc. One important aspect of this heterogeneous feature is presence of a large amount of experimental scatter in rock breakage results when considered from a quantitative perspective (Bieniawski, 1967; Liu, Roquete, Kou, & Lindqvist, 2004)."

Jan 1, 2018

By A. Chattopadhyay

Porphyry copper deposits are the most important source of copper and also gold within the orogenic belts. These deposits are characterized by low grade copper ore (0.2% to 1% copper) and host precious metals like gold, silver and platinum. A characterization tool has been developed for complex porphyry type copper-gold deposits with an aim to quantify copper and precious metal deportment. This quantification is a challenging task due to low grades and also due to the presence of both visible and invisible precious metal content in different sulphide and non-sulphide phases. Studies were initiated on various complex, ore deposits and also on samples representing various processing streams such as flotation and leach products. This paper presents the development of a systematic approach to diagnostic metallurgy using state-of-the-art optical and automated mineralogy, micro-probe, dynamic SIMS and TOF-SIMS analysis for quantification of invisible and visible precious metals content in various species. The relevance of this characterization to metallurgy will also be briefly discussed in this paper.

Sep 1, 2012

By J. Botula

In the region of the North-Bohemian Coal Basin and its environment, there, in a row of places, are found the accumulations of rocks of the sedimentary and pyro clastic origin, the preliminary composition of those shows the possibility of their being utilized for the production of ecologically sparing natural Fe- and Ti- pigments and filling substances. In many places of the regions of Chomutov and Podborany and near Bilina and also in other places, there are found raw stocks perspective for the preparation of yellow up to ochre-brown, red, brown-red and green dyes. The target of project is the verification of dispensable reserves, their mineralogical composition and chemical and physical attributes, verification of treatment ways of Fe-and Ti-pigments separation (hydrocyclones, magnetic and electrostatic separation) and their stabilization and technological painting attributes for applications in colors, coating and plastic masses, mortar mixtures etc.

Jan 1, 2014

By P. Ying

An auxiliary system connected with SEM-EDX for rapid identification of mineral and rock sample has been developed. With the aid of a synthetic spectrum library created, using the compare and match techniques provided by INCA software in EDX the mineral or rock sample names for refer-ence can be given at the same time analyzed.

Jan 1, 2014

By Jun Xiao, Yucai Zhou, Yanhong Dong, Daixiong Chen

"In this article research on the mineralogical characteristics and beneficiation process for a copper molybdenum deposits in Heilongjiang province of China has been carried out. Chemical analysis, X-ray diffraction(XRD) and Quantitative Evaluation of Minerals by SCANning electron microscopy (QEMSCAN) were conducted on samples. The bulk analysis results showed that: (1) Comprehensive recovery and separation of molybdenum & copper minerals was feasible by the flowsheet of first copper-molybdenum mixed flotation after grinding the ore samples to 61.88% -74µm and then separation flotation of copper and molybdenum after regrinding the Cu-Mo mixed concentrate to 83.26% -38 µm. The results highlight relationship between mineralogical characteristics of ore and comprehensive benefication method. (2) the combination of "" Sodium sulphide and sodium thioglycolate "" reagents are effective inhibitors of copper-molybdenum mixed concentrate, their selectivity performance is better than that of single reagent. Good flotation indexes were gained that might be directive experience to similar cases. KEYWORDS Mineralogical characteristics; molybdenite; chalcopyrite; flotation separation; inhibitor INTRODUCTION Copper and molybdenum metals are widely used in the field of metallurgy, aerospace, national defence, etc. Copper –molybdenum sulphide ore is an important source of copper and molybdenum metals. At present flotation is still the preferred methods to deal with the copper-molybdenum sulphide minerals (Yin, Z., Sun, W.et al. 2017; Ansari, A., & Pawlik, M. 2007; Hirajima, T., Miki, et al. 2017). Copper and molybdenum sulphide ores usually have the similarly good nature floatability(Miki, H., Matsuoka, et al. 2017), meanwhile most molybdenum sulphide ore has tiny grained-size(Castro, S., & Laskowski, J. S. 2015.) and closed relationship with other sulphide and gangue minerals which result in the coper and molybdenum separation problem. Base on the nature of the ore and flotation properties of the individual minerals contained in the ore, there are two methods are usually used to deal with the copper – molybdenum ore, including bulk molybdenum/copper flotation followed by separation of copper and molybdenum minerals, sequential molybdenum/ copper flotation(Liu, G. Y., Lu, et al. 2012). The latter method has the shortcoming of hard to activate the depressed copper minerals in the previous molybdenum flotation process, which may result in a large loss of the copper from concentrate, this disadvantage would avoid while adopting the former method. bulk flotation followed by separation process is to get the copper-molybdenum concentrate from raw ore first then separate flotation the molybdenum and copper minerals, benefit from the excellent floatability of both molybdenum and copper minerals ( Malhotra, D., Rowe, R. M., & Bhasin, A. K. 1986; Cao, Z., Yue, et al. 2017 ), it really easy to obtain the mixed concentrate. In order to get the high quality copper and molybdenum concentrate, the key technologies and difficulties are to realize the separation of the copper and molybdenum minerals which have the very similar natural floatability (Raghavan, S., & Hsu, L. L. 2015; Nakhaei, F. 2014)."

Jan 1, 2018

By S. Banisi, A. R. Hasankhoei, A. R. Ghasemi, E. Razi, G. A. Parsapour

At the pelletizing plant of the Gol-E-Gohar mining and industrial company two parallel dry closed grinding circuits, each with the average throughput of 300 t/h, are used. Each grinding circuit consists of a burner, a dry ball mill (6.2 m×13 m) and an air separator. The ball mill consists of a 2 m-long drying and an 11 m-long grinding chambers. The iron ore concentrate is fed to the drying chamber by a feed chute. The first 1.35 m of the drying chamber consists of 56 flights in two rows to decrease the feed moisture from 5% to below 0.2% and the rest of dryer chamber consists of deflectors to pass the material to the grinding chamber. It was found that when the feed moisture content increased from 1.3% to 3.5%, the throughput decreased by 12% (35 t/h) indicating a low performance of the dryer. Monitoring the wear rate of flights for a period of 12 months showed that the first 0.8 m length of the flights length did not experience any wear. This means that about 59% of flights do not contribute in material distribution and drying. A DEM simulation of feed chute and the drying chamber demonstrated that due to the inappropriate design of feed chute most of materials do not get in touch with the first 0.8 m length of the drying chamber. To overcome this problem, various feed chute designs with different geometries were simulated by the KMPCDEM© software. With the aim of arriving at a proper material trajectory, where the total length of the flights are used, a new feed chute was selected. The simulation results indicated that if the height of the flights is increased from 1.60 to 2.26 m and the slope is increased from 45 to 48 degrees the material arrive at the first 0.48 m of drying chamber. In this manner, the unused part of the drying chamber decreases from 59% to 36% of the length. After installation of the new feed chute, during a period of three months the throughput increased by 36 t/h.

Jan 1, 2018

By Liu Jin

"In order to reduce the dosage costs and increase the economic benefits of a full floating process in a low grade complex multi-metals W-Mo ore, based on the study of the magnetic characteristics of the main gangue minerals, a joint process by magnetic separation and floatation technology is adopted. The first step of the flowsheet is to recover magnetite by magnetic separation, then remove the magnetic gangues by high gradient magnetic separation (HGMS) and finally recycle the W-Mo from the magnetically separated tailings by flotation. Results of the lab experiment show that in comparison with the full flotation process, the floatation feed tonnage of the combined flowsheet are only 46%. The grade and the recovery of Mo in the rough Mo-Bi concentrate is similar with the original one while the WO3 in the tungsten rough concentrate is increased from 3.85% to 6.64%. The dosage costs are reduced by 53%. The new process can significantly improve the plant economically. INTRODUCTION With the continuous exploitation of scheelite resources, the easily obtained resource is gradually depleting. The ""lean, fine and complex"" scheelite resource has become the main research object of tungsten beneficiation (Lou et al., 2010) (Xiao et al., 2013). The passive situation of tungsten beneficiation can be reversed by enhancing technological innovation and investment in science and technology, changing the business philosophy and mode of production, optimizing the process structure and process conditions, improving various technical indexes and reducing production costs of reagents, which will maximize economic benefits (Jiang et al., 2012 ) (Zhu et al., 2016). A polymetallic deposit of tungsten molybdenum bismuth original process adopts a full flotation process. With the expansion of stope mining, the content of valuable metals in the ore continues to decrease, resulting in a high cost per ton of concentrate produced and a low output value (Li et al., 2015)."

Jan 1, 2018

By G. Y. Liu

The structure-activity relationships for four thionocarbamate collectors are discussed in this pa-per. These collector classes are alkyl alkyl thionocarbamates, allyl alkyl thionocarbamates, acyl al-kyl thionocarbamates and alkoxylcarbonyl alkyl thionocarbamates. The main emphasis will be to highlight the differences between collectors and their affinities for copper and iron sulfide minerals on the basis of data on electronic-structure properties obtained by vacuum and solution density functional calculations, flotation experiments and analysis measurements. The mechanism of interaction of these collectors with sulfide minerals is also discussed with perturbation theory. The theoretical study results indicate that acyl alkyl thionocarbamates and alkoxylcarbonyl alkyl thionocarbamates have best performance in these agents as flotation collectors with improved collecting power for copper sulfide minerals and excellent selectivity against iron sulfide minerals, which are consistent with the experimental data reported in the literature.

Jan 1, 2014

By T. J. Chen

Stone coal containing vanadium is an important vanadium resource in China. Roasting and oxidation of vanadium is the foundation for extracting vanadium from stone coal. This article takes stone coal vanadium ore in Tongshan of Hubei Province as research target to conduct oxidation roasting test on it by using a self-developed low-sodium compound additive and a new optimized technology. Firstly, use potentiometric titration, X diffraction, energy spectrum and electron probe to study valence change, occurrence and phase transition of vanadium during oxidation roasting process. On this basis, primary work is done on oxidation kinetics of vanadium by exploring the steps that control reaction rate during vanadium oxidation process and determining the related dynamic parameter ? apparent activation energy. The study shows that occurrence of vanadium in stone coal of Tongshan is mainly in illite, and that the result of phase transition in roasting process is disappearance of illite containing vanadium, formation of soluble vanadate and potassium feldspar, albite. It further demonstrates that oxidation process of vanadium is step by step (V(?)? V(?)? V(?)), and its oxidation reaction is controlled by diffusion kinetics whose dynamic apparent activation energy is E1=81.19kJ.mol-1.

Jan 1, 2014

By E. Burdukova

This paper explores the relationship between the rheological behaviour and mineralogical nature of three different natural ore types. Rheological measurements performed here on Platreef, Kimber-lite and Norseman-Wiluna ores are correlated with the variation of phylosilicate mineral content in these ores and the corresponding physical properties of selected phylosilicate mineral groups. Preliminary results show that the clay mineral group (talc) has the least effect on both Bingham yield stress and viscosity of the three ore slurries tested, whereas the serpentine mineral group has a dramatic increasing effect on the Bingham yield stress especially at very low solid volume fractions. The Bingham viscosity of the mineral slurry appears to be the most influenced by the presence of the swelling clay mineral group (smectite). Significant increases in either Bingham yield stress or viscosity due to elevated contents of these problematic phylosilicate mineral groups are likely to cause negative effects on the processing and manageability of mineral slurries but it is unclear which is likely to have a more detrimental effect.

Jan 1, 2014

This article takes the OMRON C200HE PLC for example introducing its application in mica-paper production equipment, including style selection, address assignment of I/O point, and programming for different function. In the transformation process of simulated data and digit data, the corresponding relations between two data and so on are also discussed.

Jan 1, 2014

The objective of the paper is to develop a material balance calculation procedure for a production unit involving two plants separated by a pipeline. The case study is taken from the oil sands industry. The raw sands are first mixed with water in a conditioning plant, then transported through a pipeline to an oil extraction plant. The combination of pipeline dimensions and nominal throughput creates a significant variable time lag between the two plants. As usual, mass balance calculations required for the evaluation of plant performance indicators, such as bitumen recovery, are inaccurate due to measurement errors. In addition, in the present case study, the time lag of the pipe creates an additional source of uncertainty, which may lead to misinterpretations and wrong decisions. Raw data are usually reconciled through steady-state mass balancing, to correct for measurement errors. However, because of the significant pipeline dynamic lag, standard steady-state daily mass balance may fail at giving an appropriate picture of the operations. The proposed solution to the data reconciliation problem includes the use of a pipeline plug-flow model in the mass balance calculation. To fulfil the industrial constraint that current data reconciliation must not modify past reconciled values, while still satisfying dynamic constraints, a daily one-step method, i.e. a unitary filter-ing horizon, is recommended. The method is illustrated with simulation data, and the results com-pared to those of standard data reconciliation methods.

Jan 1, 2014

By I. D. Ustinov, V. A. Arsentyev, A. O. Mezenin, A. M. Gerasimov

"The solid mineral raw materials of the sediment deposits containing the argillaceous variations such as the coal and the potash ores are the most multi-tonnage solid minerals mined and processed in the territory of the Russian Federation. Both types of the deposits have relatively simple geology aspects with large and medium-sized solids and as a rule, the most often, with the regular bedding which are characterized by the fitchery thickness and the internal structure. Such deposits are very difficult for mining and processing. The thermal modification of this of ores allows changing the physico-chemical properties of its components in such a way that their further processing - grinding, separation, transportation and storage can be carried out without the use of process water or significantly reduces its consumption. The purpose of the present study is to increase the efficiency of the mineral processing containing the layered aluminum silicate by reducing their negative impact on the basic dressing processes, the separation of suspended matters and the storage of waste. INTRODUCTION Many deposits of minerals are associated with the presence of layered silicates such as the argil, mudstones, argillaceous slate - highly hydrophilic mineral varieties which have the ability to swell in the water. This ability to self-disperse creates the great difficulties both for the extraction and the processing of such ores. Since the dressing of most minerals is associated with the use of wet processes, where in contact with water the swelling layered silicates are dispersed to a particle size of less than 50 µm, but due to their high hydrophylic property they form bonds that promote the formation of so-called structured suspensions having increased viscosity. In this case the structured suspensions make it difficult to perform the separation operations as the classification, the gravity and magnetic separation, the flotation and thickening and filtration operations."

Jan 1, 2018

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 Z. G. Guan

The tungsten ore in Xingluokeng belongs to a symbiotic deposit containing granite and fine vein-type wolframite and scheelite. The tungsten (WO3wolframite) accounts for only 0.233% of the raw ore, 49.60% of the , 49.60% of the scheelite and 0.8% of the tungstite. The dissemination size of the tungsten ore is extremely uneven, with 0.02~1.28 mm for scheelite and 0.01~1.28 mm for wolframite. The dissemination size of scheelite is coarser than that of wolframite. The size of 83% of scheelite and 60% of wolframite is greater than 0.04 mm. The composition of minerals in the ore is quite complex. The major metallic minerals include wolframite, scheelite and pyrite, which are followed by molybdenite, chalcopyrite, arsenopyrite, limonite, magnetite, hematite and siderite. The major gangue minerals include quartz, feldspar, muscovite, biotite and sericite, which are followed by chlorite, fluorite, apatite, dolomite, etc. According to the dissemination features and liberation degree of the tungsten minerals, the technology flowsheet of gravity concentration for coarse particles and flotation for fine particles to re-cover the low-grade and fine vein-type tungsten minerals is adopted, and -0.7 mm is selected as the processing size. The closed-circuit grinding procedure, with the combination of the rod mill and high-frequency vibration screen, is adopted. The grinding product is classified into three products of -0.7 +0.2 mm, -0.2 +0.04 mm and -0.04 mm. The gravity separation is adopted for the former two to recover tungsten minerals and the flotation is adopted for the last one. Under the condition that the raw ore contains 0.23% WO3, some technology indexes of beneficiation have been achieved. The gravity tungsten concentrate contains 63.54% WO3 with the recovery 70.11%. The slime flotation tungsten concentrate contains 35.22% WO3 with the recovery 10.48%. The composite tungsten concentrate contains 57.53% WO3 with the recovery 80.59%. Thus, the tungsten minerals are re-covered effectively.

Jan 1, 2014