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By M. Saghaeian, R. Hejazi, L. Hosseini, M. Zare

Difference between plant feed specifications in practice and design feed is a problem that most of processing plants face during operating and commissioning. This topic has led to inappropriate operating and necessitates repeating the basic tests on new feed. Gol-E-Gohar fourth concentrate line with a capacity of 2 MTPY, located in Kerman province, is the first iron concentrate plant in Iran in which High Pressure Grinding Roll (HPGR) technology has been applied. Plant&apos;s feed consists of 30% AG mills pebbles of lines 1-3 and 70% run of mine crushed ore whose Fe=51%, Fe/FeO=2.5, S=1.7% and work index=15 kWh/t. Desired concentrate has Fe>68% and S<0.7%. After the plant construction and operation commencement, providing design feed was not possible due to the crushing unit`s not being complete, therefore the plant was fed by AG mills pebbles. For predicting optimal conditions of the plant and investigating the possibility of improving the current concentrate specifications, sampling from the new feed and laboratory and pilot tests was performed again. Studies showed that new feed has Fe=41%, Fe/FeO=3.5, work index=17 kWh/t. Results of laboratory tests with Davis tube device and pilot tests run using magnetic separators with simulating the plant`s circuit showed that with current feed, achievable recovery in industrial scale would be 75.52% which compared to plant&apos;s recovery in studied period (75.94%) shows that plant operated in optimal conditions. To improve the current concentrate specifications (Fe=66%, S=1.03%, Ferec=75.94%) and for achieving desired concentrate, samples were reground to four fractions with k100= 300, 212, 100, 45 and Davis tube tests were performed on each of the samples separately. Results showed that the fraction whose k100=212 µm or k80=200 µm has respective Fe and S grades of 68.27% and 0.681% which are consistent with the desired concentrate.

Jan 1, 2014

By Caroline Olsen

The objective of this study was to achieve Zn activation control in an industrial Cu flotation circuit treating a complex Cu-Pb-Zn-Au-Ag sulphide ore at the LaRonde Division of Agnico-Eagle Mines Limited (Québec, Canada). In order to provide process relevant results, a laboratory testing methodology was developed to duplicate plant operational parameters. The action mechanisms of ZnSO4, TETA (triethylenetetramine) and NaHSO3used as depressants in order to limit ZnS activation were examined. Laboratory flotation tests with and without Cu2+ions (the ions responsible for ZnSactivation)were performed. All depressants showed equal or better flotation performance relative to baseline tests. For example, in the presence of Cu2+, the Zn recovery reduced from 32.2% to 17.5% in the copper rougher concentrate after adding ZnSO4(with a dosage of 50 g/t). TOF-SIMS analyses confirmed the relative decrease in surface Cu in the samples treated with ZnSO4and TETA. For the case with TETA, this indirectlyindicates the formation of a copper TETA complex in solution, thereby reducing available Cu for transfer. The TOF-SIMS and XPS data indicate that the addition of ZnSO4and NaHSO3resulted in an increase in surface metal hydroxides and SO3 on ZnS grains inhibiting both Cu and collector attachment.ZnSO4 was subsequently tested with success in the LaRonde concentrator. The addition of 40 g/t ZnSO4 at the conditioning (or cleaning) stage led to a 2% reduction in Zn content in the Cu and PM concentrate. Recently, improvements towards the control of copper sulfate addition in the downstream Zn flotation circuit resulted in the reduction of Cu2+ in process water. As this ?cleaner? water is recycled for Cu and PM flotation, the addition of ZnSO4 is now used only as required.

Sep 1, 2012

By M. Eraydin

Dewatering fine particles (e.g., flotation concentrates) is inefficient, energy intensive and often requires large capital and operating costs. To address these issues, a series of advanced dewatering technologies have been developed at Virginia Tech. These include novel dewatering aids and a hyperbaric centrifuge. Typically, the moistures obtained using these methods are 30% to 50% lower than those obtained using conventional methods, as demonstrated in laboratory and pilot-scale test programs. With another novel dewatering technology that has been tested in bench-scale tests, the moistures can be reduced to levels once believed to be attainable only through the use of thermal dryers. In this communication, test results obtained with these advanced dewatering technologies are presented and compared.

Jan 1, 2014

By Olga D. Khmelnitskaya

The paper presents the results of research on the development of gold recovery technology from oxidized gold-copper ore of one of deposits in Russian Federation. Gold and copper grades in the ore are 9.5 g/t and 0.4%, respectively. Various options of flowsheets, such as direct cyanidation, thiourea leaching, flotation and gravity concentration followed by subsequent cyanidation of con-centration products, were tested. It was established that gravity-cyanidation technology with filtra-tion leaching tails and the use of solutions in the recirculation is the optimal option for ore treatment. The proposed flowsheet provides gold recovery for 91.0% at NaCN consumption 2.5 kg/t ac-cording to the results of pilot testing. Ammonia-cyanide gold leaching was also examined. In comparison with cyanidation, gold leaching solution in system NaCN-NH4OH-(NH4)2CO3 allows NaCN consumption to be reduced three times from 4.5 kg to 1.5 kg per 1 ton of the ore and provides the same gold recovery. Based on the laboratory and pilot plant tests results of gravity-cyanidation technology of gold dissolution, process design criteria were developed.

Jan 1, 2014

By A. Lewis

"Thiol collectors including the xanthates and dithiophosphates have dominated the sulfide mineral processing industry for over a century. Tecflote represents completely different chemistry, without sulfur or phosphorus, and will float all sulfides and associated precious metals. As these collectors are surfactant in nature, they comprise a head group which binds to mineral surfaces, a spacer and a long hydrophobic tail. The modification of the spacer behind the head group allows for collector optimisation for different sulfide ore types and process conditions.The Tecflote surfactants have been designed to be either very selective against pyrite at the natural pH of the ore (pH 7-9) or to float all sulfides including pyrite. Depending on the ore type and the process flow sheet, the use of lime can be maintained, reduced or completely eliminated.Tecflote can be used either as a rougher collector giving good recovery and grade for the sulfide minerals or it can be used to target valuable minerals after regrind and through cleaner flotation. It has repeatedly shown to be very effective at holding onto gold through the cleaner steps. The low solubility has a significant impact on the mechanism and surface activity of the collector, resulting in a very different flotation response compared to conventional thiol collectors.INTRODUCTIONXanthate was first made in 1882 by Zeise and first used in flotation in 1924 (Keller, 2018). In 1946 Keller patented thionocarbamate as a collector, which was produced by Dow Chemicals (Keller, 2018). Later it was replaced by Dow product Z-200, while dialkylthionocarbamate was discovered by Harris et al. (Harris, 1954. No significant chemistry change to collectors has occurred since they were introduced in the mid-20th century and now that the feed grades have been declining over the last decades creating a need to find more selective collectors to extract difficult and complex ores and pushing the industry to look at other reagents."

Jan 1, 2018

By Edgar Schach, Tom Leistner, Markus Buchmann, Urs A. Peuker, Martin Rudolph, Marius Kern

"In this study, the flotation characteristics of a cassiterite-bearing fine-grained and complex skarn ore from a deposit in the Ore Mountains is investigated. The tests are performed using an oil-assisted column flotation approach to process very fine ore fractions and avoid losses of cassiterite into the tailings. First, process parameters are obtained for a finely ground artificial mixture of quartz, magnetite and cassiterite, simulating the real ore. Thereby, magnetite is used, as iron oxides can have a detrimental effect on the flotation due to a similar flotation behavior. In addition, they can act as a source of multivalent ions, which are known to reduce the concentration of collector molecules, active for flotation. Based on the results, selected parameters are further tested for cassiterite skarn ore from the Hämmerlein deposit including a pre-conditioning and a water exchange step to remove ionic contaminants. The process response is analyzed in detail by XRF (X-ray fluorescence) and MLA (mineral liberation analysis) to get a better understanding of the behavior of the single ore components. Sulfosuccinamate type surfactant is utilized as the collector, emulsifier and to reduce the froth destabilization through nonpolar oil. Sodium hexafluorosilicate is added as the depressant.INTRODUCTION Complex skarn ore deposits containing considerable amounts of tin, indium and other metals are located in the Ore Mountains region (Germany). In this study, a skarn ore from the Hämmerlein deposit was investigated. Due to the finely disseminated nature of this ore, past beneficiation attempts failed, and the ore was considered economically unprocessable. However, increasing resource demands brings it back into focus (Schuppan and Hiller, 2012). Cassiterite is the main tin bearing mineral in this ore. Due to its high density (~7 g/cm³), the predominant beneficiation technique is gravity separation (Angadi et al., 2015). Tin ores from disseminated deposits have to be extensively comminuted to reach a sufficient degree of liberation of the cassiterite grains. As cassiterite has a brittle nature, it is preferable enriched in small size fractions, during comminution (Angadi et al., 2015). Gravity separation methods become more and more inefficient for particles smaller 40 µm (Falcon, 1982), causing tin losses of up to 30 to 40wt.-% (Angadi et al., 2015). Thus, froth flotation of cassiterite was first utilized in Altenberg in the Ore Mountains as an additional processing step to reduce the cassiterite losses into gravity separation tailings (Bulatovic, 2010; Gruner and Bilsing, 1992). Although the reported lower size limits for an effective cassiterite floatability are often smaller compared to other minerals, the flotation performance is in general decreasing for small particle sizes < 10 µm (Trahar and Warren, 1976). Therefore, different approaches have been already investigated for the flotation of such fine cassiterite fractions to overcome the limits in particle size. Bilgen et al. investigated the shear flocculation of cassiterite with a sulfosuccinamate collector (Bilgen et al., 1994). The agglomeration flotation of cassiterite has been tested by Schubert et al. (Schubert et al., 1966), in which the selective coagulation of fine hydrophobic particles and oil droplets leads to an increase of the effective particle size. Recently, Leistner at al. investigated an oil-assisted agglomeration flotation for ultra-fine cassiterite ore fractions (Leistner et al., 2016). However, a poor response due to an increased concentration of Fe and Ca ions, unselective oil/gangue coagulation and froth instability, that is caused by oil/frother interactions, has been observed."

Jan 1, 2018

By Rogerio de Oliveira Baldoino

The sedimentary phosphate deposit of Bayovar is located in a very dry area of Peru (desert of Sechura) which is close to the Pacific coast. It consists of seven superposed layers (beds) which call for different strategy of concentration. The ore is composed predominantly by apatite (84% in weight basis) plus silicates (16%). In order to treat the ore from beds 1-5, a suitable standard mineral processing circuit (composed of scrubbing, screening, attrition and desliming) started its operation in 2010, yielding 3.9 Mt/y of concentrate assaying 30% of P2O5. To concentrate phosphate from beds 6-7, however, it is necessary to carry out further cationic reverse flotation of silicates, at neutral pH, to yield a concentrate, which meets market specification. The new process was developed in laboratory scale by a collaborative program conducted between Vale S.A. and the University of Sao Paulo-Brazil. A pilot plant has currently been running in order to confirm the very good results yielded in laboratory scale. The successful accomplishment of flotation in the current beneficiation flow sheet will enable a capacity expansion of 1.9 Mt/y of concentrate assaying 30% of P2O5. As water resources are very limited at Bayovar region, all the steps of concentration, including flotation, must be performed with seawater, which must be continuously recycled. This work addresses some relevant topics related to the use of seawater in the reverse cationic flotation of silicates: influence of temperature, water quality (ageing and chemical composition) and collector type (ether amine versus condensate). The results indicate that condensate amine works better than ether amine. If flotation is conducted with the former collector, the separation process is able to tolerate changes in temperature (25-40oC) and water quality caused by either evaporation or ageing. Keywords: phosphate, flotation, seawater, temperature, ageing, concentration, salt

Sep 1, 2012

By Mark Viduetsky, Dmitry Tropnikov, Victor Maltsev, Victor Sokolov, Alexander Purgin, Andrey Panshin, Alexander Bondarev, Pavel Kozlov, Igor Garifulin

"CFM pneumatic flotation machines were developed by the Institute of Materials Science and Metallurgy, UrFU, and designed to concentrate a wide range of ores and man-made materials. A key principle of work is pulp preparation for flotation owing to its saturation with finely dispersed air bubbles in ejectors. CFM plants, their advantages over domestic and foreign prototypes are described. CFM pneumatic machine commercialization resulted in: - machines demonstrated high floatation efficiency for concentration of rebellious pyrrhotite-bearing copper and zink by simultaneous floatation of copper and zinc minerals and depressing pyrrhotite. CFM- 1400M introduction allowed simplifying and optimizing the flow chart and increase extraction of copper - by 5.4%, zinc - by 9.5% and improve the quality of copper (3% Cu) and zinc (1.5% Zn) concentrates. - due to copper pyrite concentration by using CFM 1400M overall copper recovery was increased by 1.5 %, and the content of copper in concentrates - by1%. According to the results of pilot runs, the use of CFM flotation for zinc cakes concentration strongly intensifies the process of cleaning water-soluble metal compounds and, thereby, increases the efficiency of flotation concentration and subsequent milling. CFM machines showed higher specific capacity (three or four times higher) (both for feed material and froth product). Specifically, CFM froth product has higher content of useful elements (silver by 0.014 % abs. zinc by 0.31% abs.) and better extraction (silver by 6.08% abs., zinc by 0.31% abs., and sulfide sulfur by 13.7 %). Sulfide sulfur content of CFM product chamber as compared to mechanical flotation machines is two times lower (0.48 vs. 0.95 %)."

Jan 1, 2014

By Luo Chuansheng, Rao Jinshan, He Xiaojuan, Qiu Xianyang

The dissemination grain size of chalcopyrite is really fine and refractory grains of -10µm take up 25%. Dissemination relationships among chalcopyrite with pyrrhotite, pyrite, marcasite and gangue minerals are complex. Content of Fe-bearing sulphides and Fe-bearing silicates associated with chalcopyrite reaches to about 60%. It’s hard to liberate and separate chalcopyrite from pyrrhotite. The above four are the main mineral processing difficulties of a copper ore in China. Because of the similar floatability and poor separation effect between chalcopyrite and pyrrhotite, enhanced depressing and collecting method was adopted in Cu-S separation before 2009. Due to the different magnetism among chalcopyrite, pyrrhotite and Fe-bearing silicates, magnetic separation is creatively induced in Cu floatation circuit based on the in-depth study of process mineralogy. High efficiency and new Cu-S separation technology of "Rough concentrate regrinding-pyrrhotite rejecting from middling by magnetic separation" and a highly selective copper minerals collector YL-1 had been developed to solve the difficult problems of refractory Cu-S separation, which are heavy load of middling recirculation,difficulty in upgrading Cu concentrate, unstable locked cycle and low separation indexes. And grade of Cu concentrate has been increased by 3.33%, recovery of Cu and Ag has been increased by 19.32% and 10% respectively.

Jan 1, 2014

By F. Waanders, Campbell. Q.

This project was specifically aimed at using increased acceleration separation, as a method to optimize the recovery of gold in an ore body mainly consisting of hematite where the specific gravity of gold is much higher in comparison to the carrying material, making it possible to separate gold from other materials like silica, base metals and their oxides, usually associated with gravitation-gold-recovery-processes. The ore body investigated in this project originated from a gold reef containing a large portion of gold locked inside the gold pyrite complexes, which was mined and treated in a conventional way. After gold extraction, the material was pumped to a tailings dump and the gold extraction of this dump was investigated. In order to assemble a concentration plant setup, the head grade and size distribution of the dump and bulk sample were determined accurately. Chemical analysis and mineralogical examination were done on a representative sample of the bulk sample, to determine the chemical composition of the material. The results obtained thereof were evaluated and a pilot plant was configured. A bulk sample was processed in the pilot plant and the results were evaluated. The results at optimum acceleration forces applied, resulted in a recovery of 5% of the mass, with a gold concentrate of 90 g/t Au, which represented 58% recovery of the gold. The hematite with high specific gravity as main phase positively influenced the high acceleration separation process. It proved that if the specific gravity of particles in a suspension were increased, high acceleration separation could be applied effectively to separate much smaller particle sizes.

Jan 1, 2014

By S. Cagirici, Z. Öztürk, A. Hassanzadeh

Many flotation kinetics models have been studied in the literature. Their applicability was extensively investigated and argued in detail. However, model selection criteria were not adequately discussed from the statistical points of view. In this investigation, the kinetic behavior of a complex copper sulfide ore was studied in a mechanical Denver flotation cell focusing on flotation kinetics of chalcopyrite, pyrite and molybdenite. Different flotation kinetics models including nine common empirical models and four mathematical models namely Hill, Chapman (Sigmodial function), single rectangular (Hyperbola equation) and exponential were applied to the experimental data. In addition to assessment of the goodness of fit criterion for each model, a factor of model complexity was considered using information criteria (IC) (i.e. Bayesian information (BIC), low of iterated logarithm (LILC) and Akaike information (AIC) indices). The obtained results showed that the IC indices could simply manifest the best-fitted model to the experimental data. Whereas, the coefficient of determination values (R2) were relatively same for all models. By taking the R2 and model complexity criteria into account, the exponential model was chosen as the best representative mathematical model to demonstrate chalcopyrite kinetic behavior. However, Chapman model was selected as the best one for the flotation of pyrite and molybdenite. In case of the common first-order flotation kinetics models, fast and slow flotation kinetic model (Kelsall) was reasonably fitted the best to the given data of chalcopyrite. However, the gas/solid kinetics adsorption model was chosen as the best-fitted one for pyrite and molybdenite. Furthermore, it was found that mathematical models represent better results in association with flotation kinetic behavior of chalcopyrite, pyrite and molybdenite due to the consideration of more parameters in modeling. Finally, it was concluded that the IC indices must be applied to the process of model selection due to consideration of goodness of fit, complexity of a model and model consistency.

Jan 1, 2018

By Shibayama A.

In this research, an investigation of a process to selectively remove arsenic (As) and antimony (Sb) contained in enargite (Cu3AsS4) and tennantite/tetrahedrite ((Cu1,Fe)12(As,Sb)4S13) copper resources by leaching and precipitation were performed. First in the leaching process, the target was to selectively leach As/Sb and reduce their contents in the ores to <0.5 wt% followed by a second process to remove As/Sb dissolved in solution for environmental control. Enargite ore contained 3.65 wt% As while tennantite/tetrahedrite ore contained 5.40 wt% As and 4.85 wt% Sb. Alkaline leaching using NaHS?NaOH media at varying experimental conditions (temperature, pulp density, time, reagent concentrations) were performed to investigate the process. As/Sb precipitation studies were also performed to control high As/Sb concentrations in leach solutions by using elemental sulphur (So). Results from leaching studies showed that over 99% of As/Sb in the impure copper ores can be removed and their contents in the final solids reduced to <0.5 wt%. It was also established that such very high leaching could be achieved at 80°C within 1?6 h and pulp densities as high as 1000 g/L. The study also revealed that leach solution containing very high As/Sb concentration ranging from 5 to 20 g/L can be used as fresh solution for leaching of fresh ores until concentrations exceeding 35 g/L when it was sent to precipitation circuit for As/Sb removal. Addition of elemental sulphur at 90°C and fast cooling to 30°C allowed over 60% As/Sb precipitation and removed as Na3AsS4 and solution with reduced As/Sb concentration to leaching circuit for continuous use. Keywords: Enargite; Tennantite/Tetrahedrite; Selective Leaching; Precipitation; Sodium hydrosulfide;

Sep 1, 2012

By Shailin Lalloo, Allan B. Nesbitt

A fluidised bed was exposed to high gravities by placing it in a centrifuge. The bed expansions were measured and compared with the bed expansions predicted by the Richardson-Zaki model (1954). The Richardson-Zaki model was found to predict effectively the fluidisation behaviour of defined particles in this increased gravity environment.

Jan 1, 2003

By V. Panayotov

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

Sep 1, 2012

By A. Yörükoglu

In this study, thorium mineral that found in the complex ore deposit bearing bastnasite, fluorite, and barite from Eskisehir - Sivrihisar district has been tried to extract by using HCl and H2SO4 leaching. Ore sample of approximately 3 tons taken from the ore deposit has been homogenized in order to obtain composite samples by means of ore preparation techniques for experiments. Results of chemical analyses of composite samples are as given; 40.8% CaF2, 35.35% BaSO4, 2.01 Ce, 1.5% La, 0.39% and 0.1% Th. HCl and H2SO4 leaching processes have been performed to gain Th in the ore. For determining the optimum particle size, the samples have been ground under controlled conditions and brought to the sizes of 0.065, 0.2, 1 and 3 mm. Leaching experiments have been carried out in both HCl and H2SO4 acids without roasting and roasting at the temperature of 900°C for 1, 3 and 6 h retention times for all particle sizes determined before. In the experiments, the highest Th recoveries reached are as follows; 78.74% Th when H2SO4 and 1 mm particle sized sample used, and 1 h roasting applied and 73.38% Th when HCl and 1 mm particle sized sample used, and 3 h roasting applied. Keywords: thorium, leaching, concentration, sivrihisar, roasting

Sep 1, 2012

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 C. A. Perez

Estimation of rock composition in mineral plants is important to determine rock size and grindability which in turn may improve control of the grinding process. Variations in ore grindability and size distribution directly affects mills power consumption and throughput. Therefore, it is desirable to develop new tools to estimate rock lithological composition. In this paper, a new method for lithological classification based on machine vision is proposed. The method is based on a single video camera and employs information about texture using several different spatial scales. Texture information is extracted from rock images using Gabor filters that can be adjusted at different spatial scales and orientations. After feature extraction, rock images are classified using support-vector machine (SVM). The method was tested on a database containing sub-images of 64x43 pixels of five ore types as follows: massive sulphide (MS), disseminated sulphide (DS), net textured (NT), gabbro (G), and peridotite (P). These ore types were assigned to three grindability classes: soft (MS), medium (DS and NT) and hard (G and P). The database was divided in 2 subsets for training, cross validation, and for test. Classification accuracy is compared with previously published results on a dataset with three different classes of rocks according to their hardness. Results show that our proposed method reached significant improvements (between 13%-20%) in classification accuracy relative to previously published results. Keywords: rock classification, lithological classification, grindability estimation, Gabor feature extraction, texture features

Sep 1, 2012

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 Yuhua Wang, Haitao Zheng, Guangli Zhu, Yuehao Zhao, Fushun Yu, Dongfang Lu

"Spodumene, feldspar and quartz possess similar surface properties, which make the difficult to separate spodumene from other silicate minerals by flotation. The differences of crystal structure, surface property and floatability of three minerals were studied by various technical means. The results show that Li-O and Al-O bonds in spodumene structure are not strong and easy to be broken. However, Al-O and Si-O bonds in feldspar and Si-O bond in quartz are very strong and not easy to be broken. High contents of Al and Li on spodumene surface were found by XPS measurements, which is benefit for the adsorption of anionic collector. The surface energies of main cleavage planes of spodumene (110), feldspar (001) and quartz (001) are 1.88 J/m2, 2.97J/m2 and 2.68J/m2, respectively. Hence, the exposed surfaces of feldspar and quartz are not saturated and easy to adsorb OH- ions from solution and cause the lower IEP than spodumene. The contact angles of three minerals in water are lower than 16º, suggesting that the hydrophilicity of three minerals is high, and the order of hydrophilicity is quartz > feldspar >spodumene. Correspondingly, contact angles and flotation recoveries of spodumene are bigger than that of feldspar and quartz in presence of a combined anionic and cationic collector named YOA.INTRODUCTION Flotation technology plays most important role for the recovery of spodumene from pegmatite ore. Generally, spodumene with fresh surface is easy to be floated using soap and oleic acid as collector, and the best flotation pH range is 4.0 to 9.0. However, the flotability of spodumene will become worse when its surface was polluted by slime which produced by the weathering of spodumene and other silicates (Dresler et al. 1998; Moon and Fuerstenau 2003; Menéndez et al. 2004; Yu et al. 2015). In addition, the species of Ca2+, Fe3+, Mg2+ and other metal ions in pulp can not only activate the flotation of spodumene, but also the activation of gangue minerals, and then the flotation selectivity of spodumene from gangue minerals became wores (Wang and Yu 2007; Yu et al. 2014). In order to eliminate the negative effects of slime and metal ions on the flotation of spoddumene, the desliming and long time scrubbing with the addition of NaOH and Na2CO3 were adopted. Usually, NaOH shows the activation for spodumene flotation, and the recovery of spodumene increases with the increase of NaOH dosage. Na2CO3 can eliminate the non selective activation of Ca2+, Fe3+, Mg2+ ions in the pulp, and the concentrate grade of spodumene increases with the increase of Na2CO3 dosage. Anionic collectors such as chelants, hydroxamic acid, oxidized paraffin soap, naphthenic acid and oleic acid were widely used for the flotation of spodumene (Dresler et al. 1998; Moon and Fuerstenau 2003; Menéndez et al. 2004; Yu et al. 2015)."

Jan 1, 2018

By A. Erdem

In this study, approximately 3 tons of coal sample taken from thickener entrance was prepared by mineral processing techniques and cleaned by gravity methods. The samples were classified to 5 size fractions by wet screening. The size fractions of +1mm and -1 +0.500 mm were chosen for direct consumption or cleaning by spirals. The studies by Multi Gravity Separator (MGS) were carried out with the -0.500 +0.280 mm and -0.280 +0.020 mm size fractions where the -0.020 mm fraction does not used in the experiments. The clean coal fractions of +1 mm and -1 +0.500 mm obtained by classification were subjected to lower calorific value (LCV), ash and total sulphur analyses. The results were 4168-4946 kcal/kg (on dry basis), 31.01-38.54 % ash and 4-4.27 % total sulphur content respectively. The results of the clean coal concentrated by MGS experiments in -0.500 +0.280 mm fraction was found 5918 kcal/kg LCV, 19.82 % ash content and 80.47 % combustible recovery. The results for the -0.280 +0.020mm fraction were 6004 kcal/kg LCV, 19.19 % ash content, 80.47 % combustible recovery respectively. With reference to the above experimental results, pre-economic evaluation studies were completed and decided to construct an additional unit to the plant. Capacity of the constructed unit is 75.000 t/y which corresponds $ 2.600.000 income in a year. General Directorate of Turkish Coal Enterprises concluded to apply this project to the other coal washing plants in the country due to the successful results of this study and laboratory studies has been going on by MTA currently. Keywords: coal tailings, mgs, recovery, washery, pre-economic evaluation

Sep 1, 2012