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By Carlos Collado, Oscar Jimenez, Francisco Diaz, Rodrigo Bernal, Carmen Salinas

Water is a strategic resource for the mining industry. The rising cost has forced the industry to become more efficient in their use, reduce and recycle waste as much as possible. The importance of having a reliable water balance is increasingly urgent, because the resource is scarce and vital to the sustainability of human life, animal and ecosystem. Conventional flowmeters can suffer disruptions or failures by corrosion, fouling, presence of elements in the water as acids, salts and suspended solids. Use of radiotracers is advantageous to determine fluid velocities that are transported through pipes, as they are detectable by means of the radiation emitted, the same way for surface flows, such as natural watercourses and drainage from tailings dams, fluorescent tracers are those with the best advantages. For velocity measurements in pipelines a Iodine-131 tracer solution, around 5 mCi was injected into the suction of the pumps. The technique used is the transit time between detectors at known distances and determining the time of passage of the radiotracer. For measurements of surface fluxes using the fluorescent tracer Rhodamine WT by tracer dilution technique, injecting the tracer at a constant rate and measuring it known concentration diluted downstream. This work has been performed in Minera Los Pelambres, located in the IV Region of Chile, measuring the flow rate pumped from 12 pools with 1-7 pumps operating, obtaining speeds of 0.56 to 2.62 m/s. In pumping wells three facilities have been measured at speeds of 0.111-0.223 m/s, plus six surface flows with flow rates of 8.45 to 357.2 L/s.

Jan 1, 2014

By Hideki Yamamoto, Norihiro Murayama, Junji Shibata

"Hydrothermal syntheses of zeolite from coal fly ash were carried out by using various alkali sources. NaOH, Na2CO3, KOH, K2CO3 and their mixture were applied for an alkali source of zeolite synthesis. The various physical properties of the reaction product, such as the kind of zeolite crystal, cation exchange property, surface texture and so on, was investigated through the zeolitization reaction from fly ash. The mechanism of zeolitization reaction was considered from these results.Zeolite P(NaP), hydroxysodalite(HS) and chabazite(K-CHA) are synthesized as zeolite species, when fly ash hydrothermally reacts in various alkali solutions at 393K. The crystal structure of the products obtained in NaOH solutions for 3h changes from NaP to HS with increasing NaOH concentration. K-CHA formation in KOH solutions needs 10h of reaction time, which is longer than that of NaP or HS.Dissolution, condensation and crystallization steps exist in the zeolitization reaction from fly ash, and the particle surface of products remarkably changes in each step. The fly ash is spherical particles and the surface is smooth(Fig.1(a)). The dissolution of fly ash occurs in the heating process until 393K, and the particle surface changes like unevenness(Fig.1(b)). In case of NaP, the condensation of aluminosilicate ions takes place at 373-393K, and the amorphous aluminosilicate is deposit on the unreacted fly ash(Fig.1(c)). The crystallization of NaP already begins to occur at 393K and the surface is covered with fine zeolite crystals for 24h(Fig.1(d))."

Jan 1, 2003

By L. J. Li, Q. Zhang, Z. H. Liu, Q. L. Chen

Phosphogypsum (PG) is an industrial waste from the production of phosphate fertilizer. The utilization of PG is restricted by its containing impurities, and the pretreatment is crucial to eliminating these impurities. The purpose of the research is to recycle pretreatment PG as raw material for preparation of a-calcium sulfate hemihydrate (a-HH) by salt solution method. The effect of pretreatment technologies of lime neutralization and washing on dehydration rate, phase composition and morphology evolution of a-HH were investigated by crystal water content analysis, XRD and SEM. Furthermore, the change of pH in the reaction process was also discussed in detail. The results show that the dehydration rate of conversion PG to a-HH increases and the corresponding induction period, crystal growth period and total time decrease with an increase in CaO dosage from 0.2% to 0.6%. In the presence of 0.4% CaO or 0.6% CaO, the a-HH presents a long hexagon-prism-shaped crystal with a purity of 94%, and the end face of a-HH presents polyhedron. The system pH sharply decreases with the reaction time under different CaO dosage, and the pH variation from the initial alkalinity to acidity. The main reason is that as the co-crystallized phosphorus dissolved and released HPO42-, and with the pH decreasing to 4-5, HPO42- transformed into H2PO4-. The dehydration rate of washing PG is lower than that of original PG. Compared with the original PG, a-HH with the larger width and smaller aspect ratio can be achieved from the pretreatment PG subjected to lime neutralization and washing, meanwhile the crystal surface is more smooth.

Jan 1, 2018

By P. Saha, D. M. Chakrabarti, J. P. Srivastavam, P. Bhattacharyya, B. R. Sarkhel, Banerjeem B.

The phosphatic soil deposit at the Sukulu hills region of Uganda contains 10-12% P2O5. The ore occurs as a naturally degraded soil deposit containing a very sticky slime composed of earthy lateritic minerals. Desliming of the ore, which is imperative for its further beneficiation, is an extremely difficult hurdle to overcome for its commercial utilisation. Earlier studies on samples of the same ore at the National Metallurgical Laboratory (NML) and abroad established that it was necessary to pass the ore through a series of three stages of desliming, each consisting of a step of vigorous attrition scrubbing followed by hydrocycloning, for effective elimination of the sticky slime. In the present paper an innovative use of the very strong shear-fields created inside a hydrocyclone has been discussed, that could effectively eliminate the stubborn slime from the ore. The standard design of the hydrocyclone was suitably modified to increase the shearing effect on the slurry. It was possible, through the application of this modified hydrocyclone, to not only totally eliminate all the stages of attrition scrubbing, but also to reduce the subsequent stages of hydrocycloning. The modified technique yielded a concentrate of grade comparable to that obtained by employing the earlier more complex flow sheet, with about 25% higher P2O5 recovery.

Jan 1, 2003

By Peter J. Coghill, Michael J. Millen, Brian D. Sowerby, Kirby D. Cunningham

"CSIRO Minerals has developed an integrated particle size analysis system for deployment in the grinding circuits of minerals processing plants. The system includes a slurry sampling station, a de-aeration tank and an ultrasonic particle size analyser (UltraPS). The UltraPS analyser is based on ultrasonic attenuation and velocity spectrometry combined with gamma-ray transmission. The technology is applicable to the measurement of 0.1 to 1000 µm diameter particles directly in highly concentrated slurries without dilution. The analyser incorporates a gamma-ray transmission gauge utilising an ultra-low activity radioisotope source that is exempt from International Atomic Energy Agency (IAEA) and Australian licensing regulations.The UltraPS analyser consists of a submersible probe and an electronics package. The probe can be immersed directly in a minerals processing plant stream if the aeration level in the stream is low. For applications where a de-aerated slurry stream is not available, CSIRO has developed an integrated system that includes a sampling station, a de-aeration tank as well as the UltraPS analyser itself. A schematic drawing of the continuous flow gravity fed system is shown in the figure. A commercial sampling system is used to supply slurry at a rate of 2.5m3/hr to the deaeration tank. The aim of the de-aeration tank is to remove air in the slurry to a level where meaningful particle size measurements can be performed. The tank volume is about 0.2m3, thus giving the slurry flowing through the tank a mean residence time in the tank of approximately five minutes. Tests at a number of concentrators have shown that five minutes is sufficient time for slurry to de-aerate to a level where ultrasonic measurements can be made. The tank incorporates an ultrasonic level detection sensor and flow to the deaeration tank is adjusted based on the level of slurry within the tank. The UltraPS probe is mounted through a flanged section of duct, which is bolted to the bottom of the de-aeration tank."

Jan 1, 2003

By H. U. R. Kattentidt, W. L. Dalmijn, T. P. R. de Jong

"Research at Delft University aims to develop systems for automatic inspection and sorting of bulk solids which can be applied in the mineral processing and recycling industry. A recent development includes an automatic sorter based on airport X-ray inspection technology. It is designed for continuous inspection and sorting of 5 - 150 mm solids and has as objective a cost effective and high capacity on-line inspection and sorting.Dual energy X-ray transmission imaging provides instantaneous data about particle dimensions and the atomic number distribution. The system provides particle thickness and average atomic number on a pixel level at scanning speeds between 250 and 500 lines / second. With existing X-ray components inspection speeds between 0.5 and 2 m/s at approximately 1-meter inspection width and between 1.5*1.5 and 5*5 mm2 sensor resolution can be achieved.Imaging algorithms that were developed by Delft University for applications in recycling, waste cleaning and other applications acquire shape, size distribution and composition data. Currently on-line processing of a product at a rate of around 200 particles per second is possible using a standard PC. Experiments with a demonstration X-ray unit indicated good potential for various applications in solids processing. They include continuous inspection of organic content, sulphur, chlorine, ash, and metal distribution. It is applicable for coal cleaning, ore sorting, industrial minerals, and various applications within waste cleaning and recycling. Dual energy X-ray transmission fundamentals and recent progress in system development and applications are presented."

Jan 1, 2003

By C. Suazo, O. Alruiz, A. Hofmann, I. Fernandez, W. Kracht, N. Mora

"Compañía Minera Dona Inés de Collahuasi, CMDIC, has developed a methodology that allows estimating economic benefit taking into account the variability of the ore characteristics (geometallurgical models). Currently, the grinding model allows predicting the treatment capacity, specific power and balls consumption for short, mid and long term planning. An alternative impact test to estimate SAG grinding parameters is presented in order to bring the geometallurgical model from mid and long term to short term planning. The impact test uses either the production bench or geological drill holes. This allows Geologists increasing the density of grinding parameters within the block model improving the geostatistical estimation of the rest of the blocks. The test requires a small amount of sample to obtain the requires grinding parameters for the grinding model in comparison to the well known grinding test and also, it can be performed easily and quickly by the geology team of a mining company. The alternative grinding test is presented in this article. The grinding model using this short grinding test has been successfully used in mining planning and the accuracy of the model is demonstrated using monthly average instantaneous throughput from January 2013 to March 2014, showing an average relative error of 5% and an R2 value of 0.9. Also, 111 individual variability samples were divided into two each, in order to compare the estimated grinding parameters based on this alternative test to the grinding parameters obtained from the conventional grinding test. The comparison shows an average relative error of 4% and an R2 value of 0.9."

Jan 1, 2014

By Olgac Kangal, Ali Guney, Gulay Bulut, Guven Onal, Ayhan A. Sirkeci

"This paper presents the results of pilot scale column flotation applications in the two different copper concentrators in Turkey. The Kure Copper Plant is located in the Black Sea Region of Turkey. There are four types of copper ore formations in the Küre region. The so called disseminated ore contains 60-70% siliceous minerals of which Cu content is around 0.5%. The remaining three ore types in the region are generally named “massive ores”. The massive ores may assay up to 15% Cu and the major gangue mineral is pyrite associated with relatively little amount of silica. The main copper mineral in all formations is chalcopyrite that is generally finely disseminated in pyrite therefore satisfactory liberation requires ultrafine grinding especially in the case of the massive ores.A blend of these ores, containing 1.7% Cu on average, is fed to the Küre Copper Concentrator whose capacity is 1.2 million tons of ore annually. Unfortunately, today the plant operates below its designed capacity treating around 550,000 tons of ore per year. The design of Küre Copper Plant was undertaken by the Outokumpu Company of Finland. The operation of the plant is almost automatically controlled and the feed, the concentrate and the tailings are assayed for copper by on-line XRF analysers. The particle size analysers are also connected to the grinding circuit. The feed material is ground almost below 50 µm and the flotation is conducted around pH 11.2 using Hostaflot X-231 (a trade mark of Hoechst) as the collector. Apart from the rougher cells, there are four cleaning stages after which the final concentrate is produced. A very fine grinding requirement for an adequate level of liberation makes successful flotation difficult and consequently the final concentrate can hardly assay above 16% Cu leading to copper losses around 20% in the tailings that contains 0.3% copper."

Jan 1, 2003

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 C. Dorfling, P. Albertyn

"Hydrometallurgical recovery of metals from printed circuit board waste typically involves multiple leaching stages; conventionally, base metals dissolution by mineral acids is followed by leaching of gold using cyanide or halide based lixiviants. Alternative lixiviants for gold leaching can reduce the potential hazardous nature of operation. The aim of this study was to investigate the feasibility of using thiosulfate as lixiviant for gold leaching from waste printed circuit boards. Batch leaching tests were performed to investigate the effects of thiosulfate concentration (0.1 – 0.2 M), ammonium concentration (0.2 – 0.4 M), pH (9 – 10.5), solids content (25 – 50 g/L), and copper content of the feed material on the gold leaching behavior. The concentrations of metals in liquid samples were determined by means of inductively coupled plasma atomic emission spectroscopy, and high pressure liquid chromatography was used to evaluate thiosulfate degradation/consumption. The ratio of thiosulfate to ammonia was found to have the most significant effect on the extent of gold leaching, with the best gold extraction observed at a ratio of 0.5 mole thiosulfate per mole ammonia. The lower pH range of 9 – 9.5 generally produced better Au leaching, while low copper extraction during the first stage base metal leaching stages has proven to be detrimental to gold leaching from the residue. Thiosulfate degradation/consumption increased and gold extraction decreased with higher residual copper in the feed. An optimum gold extraction of 78.5 % from base metal leach residue was achieved with 0.2 M thiosulfate, 0.4 M ammonia, and 0.02 M cupric ions at 25°C in the pH range 9 – 9.5.INTRODUCTION Background and Aim The recycling of e-waste is becoming increasingly important as it provides certain economic benefits apart from the obvious environmental benefits. Printed circuit boards (PCBs) are found in most forms of e-waste and contain especially high concentrations of base and precious metals. Hydrometallurgical processes are often considered for the recovery of valuable metals from PCBs; these processes normally implement several leaching stages to selectively recover certain metals."

Jan 1, 2018

By Xin Tan, Changgen Li

"The floatabiliby of chelate CF-collector for zinc and lead oxide minerals was described. It found that CF-collector possesses a strong collection power for zinc and lead oxide minerals as well as excellent selectivity for the gangue minerals. A new technology of zinc and lead oxide flotation with CF reagent as a collector, lead nitrite as a activator and water glass as modifier was presented. The mechanism of interaction between reagents and minerals was discussed by means of adsorption, zeta potential, nephelometric titration, XPS and infrared spectra was investigated.CF-collector, newly developed by Beijing General Research Institute of Mining and Metallurgy (BGRIMM), is a kind of chelate collector for the flotation of oxide minerals. CF-collector mainly consists of N-nitroso-N-phenylhydroxylamine ammonium salt. The CF-collector was used for the selective flotation of scheelite in the Chinese plants.In this paper, the effects of the dosages of activator Pb(NO3)2 and depressants (sodium hexametaphosphate, polyphosphate, water glass and salified water glass) on the floatabilities of several minerals (cerussite, smithsonite, calcite, dolomite, limonite and quartz) as well as their mixtures were investigated in the presence of the CF-collector. The results show that the CF-collector is the excellent collector for selective flotation of cerussite and smithsonite. Pb(NO3)2 can selectively activate the oxides of lead and zinc resulting in a decrease of the CF-collector dosage. The sodium hexametaphosphate plus the salified water glass are the effective depressants for gangue minerals. Based on the above facts, the “CF process”, namely, used CF reagent as a collector, Pb (NO3) 2 as a activator, and sodium hexametaphosphate plus salified water glass as depressant, has been used for the flotation separation of the lead oxide mixtures or zinc oxide mixtures without disliking and heating. For the feed lead assaying 19.07%, the lead concentrate assaying 49.81% Pb at the recovery 80.69% was obtained after one step flotation. For the feed zinc assaying 12.85%, a zinc concentrate of assaying 29.56% Zn at recovery 83.87% was obtained."

Jan 1, 2003

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

"Over the past few years CSIRO Minerals has been conducting research aimed at developing a continuous copper converting process to replace the traditional two-stage batch process that is currently used to produce blister copper. Sirosmelt type reactors and calcium-ferrite (lime-based) slags have been used rather than the conventional iron silicate slags. The research work has now been extended to focus on the development of a single-stage continuous copper making process (so-called direct copper converting) in which copper concentrates are converted to low-sulfur blister copper in a single step, that is the smelting and converting steps occur in one continuous stage.One feature of the direct copper converting process is that copper losses to the slag phase can be high, meaning the slag produced needs to be treated to recover the copper value. The feasibility of the whole copper making process will most likely depend on the efficiency and cost effectiveness of the copper recovery stage. Flotation has been identified as the most economical option to recover the copper from direct copper converting slags where it has been estimated that a copper recovery from the slag of above 90% is required to make the overall single stage copper making process economically feasible.A limited series of laboratory batch flotation tests was conducted on a suite of calcium-ferrite-based slags made in laboratory smelting trials. Copper sulfide concentrates from MIM were used as feed-stocks. The objectives of this work were to develop a robust flotation process to maximise the copper recovery and iron rejection from slags made in the process, and to establish operating conditions for future continuous flotation pilot plant trials.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 and the CaO/Fe ratio varied between 0.30 and 0.54. In general, the higher the copper content the more oxidised the slag. 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. Reagent additions were typically 200 g/t KEX and 500 g/t Na2S.9H2O and the grind size was generally a P80 of about 40 µm."

Jan 1, 2003

By C. L. Lin

Previous research on particle damage during high pressure grinding roll (HPGR) crushing was limited by the resolution of point-projection micro CT systems and the extent of damage was described by the percent of particles cracked for different crushing methods as determined at a voxel resolution of 40 microns. With recent advances in X-ray optics, the resolution afforded with point-projection micro CT systems has been improved by at least an order of magnitude using a lens-based X-ray micro CT system. With the use of X-ray lens imaging the resolution is determined by the lens rather than by the spot size of the micro-focus X-ray source. The high resolution X-ray micro CT (HRXMT) system made by Xradia with a voxel resolution of less than 1 micron is now being used for the characterization of particle damage during HPGR crushing. In this way, the direct 3-D determination of the internal damage of crushed ore particles which vary from 40 mm down to 100 microns in size is being studied at a voxel resolution of a few microns. The number, size, and shape of cracks can now be measured at this improved resolution and for the first time the specific crack surface area distributions of the crushed product can be calculated from HRXMT images using specially designed software tools for 3D analysis. Results from selected crushing operations are presented. Copper ore particles from HPGR crushing at high pressure (6 N/mm2) are severely damaged having crack specific surface areas as great as 10 mm2/mm3. Keywords: HPGR, particle damage, X-ray micro tomography, crack surface area

Sep 1, 2012

By Regis Texeira, Leopoldo Gutierrez, Vicente Hernandez, Caludia Vidal, Lina Uribe

"Molybdenum is currently the most important by-product of Chile´s copper mining industry. This mineral has gained international relevance over the years, because of its price. At mining deposits in Chile, copper and molybdenum are extracted together and then separated in a concentration process consisting of several flotation steps using sodium hydrosulfide (NaHS) as copper depressant. However, there are environmental issues related to the potential emisions of toxic hydrosulfide gas and bad odours, which affects nearby communities and suppose a risk for operators. This problem has led to the need for development of new technologies to eliminate the use of NaHS in the process of selective flotation of Cu and Mo minerals. In this work, results of flotation tests in which lignin derivatives reagents are used to obtain the selective separation of Cu and Mo minerals by flotation are presented. In this technology molybdenite is depressed by the action of lignin derivatives (lignosulfonates), while copper sulfides remain hydrophobic and float. The use of lignin derivatives in the selective flotation of copper molybdenum minerals represents a non-toxic alternative to NaHS.INTRODUCTIONSeparation of copper sulfides from molybdenite is commonly achieved through a two-stage flotation process. In the first stage the collective flotation of copper sulfides and molybdenite is promoted to generate a collective concentrate which is then treated in a second stage with sodium hydrosulfide (2.5-3.0 kg/t NaSH) to depress copper sulfides flotation without affecting the flotation of molybdenite. As a result a copper concentrate with 30-35% Cu, and a molybdenum concentrate with 50-52% Mo are generated. Although this process is widely used at industrial level, environmental issues related to the risk of H2S generation and bad odors have always been a concern. Thus there is a need in the industry for a more environmentally friendly process to separate copper sulfides and molybdenite."

Jan 1, 2018

By K. Sasaki, H. Miki, Y. Muta, T. Hirajima, G. P. W. Suyantara

"Selective flotation of chalcopyrite and molybdenite is continuing problem since its selectivity is not perfect and current NaHS process needs closed system flotation plant because of hydrogen sulfide emission on acidic condition. Some alternative additives investigated as chalcopyrite became hydrophilic and molybdenite kept hydrophobic. XPS analysis indicated chalcopyrite became hydrophilic since iron and copper sulphate, hydroxide was covered on the surface whereas molybdenite kept hydrophobic since only molybdenite can be seen. Flotation experiments of chalcopyrite and molybdenite mixture sample indicated excellent separation that 95% of molybdenite can be recovered as froth and 93% of chalcopyrite can be recovered as sink. Flotation experiments with real bulk flotation concentration indicated similar separation result compared with current NaHS additive system. Possible mechanisms and detailed phenomenon were investigated based on the surface analysis and thermodynamic calculation etc. this alternative additive is easy to handle compared with NaHS since it does not emit effluent gas and commercially preferable.INTRODUCTIONMany copper mine often associates molybdenite (MoS2) as trace mineral. It is important to separate molybdenite and copper mineral efficiently since molybdenite is valuable mineral and also contamination of molybdenite into copper concentrate decrease concentrate value. Conventional separation of copper sulfide and molybdenite has been achieved with sulfide bulk flotation followed by selective molybdenite flotation with NaHS addition since NaHS works as a depressant for copper sulfide (Bulatovic, 2010; Ansari and Pawlik, 2007b). However, this current process shows imperfect molybdenite recovery (Liu et al., 2012, Zanin et al., 2009) and depressant such as NaHS needs to be used under specific flotation plant since it might be environmental contaminant (Ansari and Pawlik, 2007a). there are still unknown on the effect of depressant NaHS, but one of the theory is NaHS reductant effect. It is worth to investigate the effect of other reductant as depressant on chalcopyrite. On this research, typical reductant, Na2SO3 are added to utilize for selective flotation of chalcopyrite and molybdenite."

Jan 1, 2018

By F. L. Mentg

The utilization of acid leaching tailings was studied. It is obvious that the Fe in acid leaching tailings can be leached into solutions as Fe2+. In this paper, the influence of conditions, including the concentration of ferrous sulfate, pH, temperature and velocity of air flow was investigated. The optimum conditions for preparing seed crystal are as follows: the concentration of ferrous sulfate is 20 g/L; pH is 9.5~10; reaction temperature is 10~20?; velocity of air flow is 0.4m3/h. The optimum conditions for preparing iron oxide red are as follows: the concentration of ferrous sulfate is 5 g/L; reaction temperature is 80~85?; velocity of air flow is 0.13m3/h. At last, the product of iron oxide red could be prepared by ammonia method.

Jan 1, 2014

By Y. H. Tang, H. Y. Zhang

"Tailings from lead-zinc mine contain plenty of dolomite and low-grade fine-grained barite, which is difficult to enrich by mineral processing. Since the density of barite and dolomite has some differences due to the fine dissemination size in a complex sedimentary, the gravity separation tests were conducted with no results obtained. The magnetic separation was conducted to remove the magnetic gangues, and the results were not satisfactory because of the gangue minerals such as limonite and others were low in content. The results of floatation tests showed that the mineral separation of barite and dolomite is quite difficult because that they have similar mineral floatability. Experimental research shows that the good results can be obtained by “desulfurizationbarite floatation” processing with the new depressant TS. The barite concentrate containing BaSO4 91.27%, BaSO4recovery is 70.25% when sodium dodecyl sulfonate serves as a collector and TS as the depressant.INTRODUCTIONBarite is an important mineral material. Often associated with other minerals, it usually needs to be separated by mineral processing in order to meet the application requirements. The main barite beneficiation methods include hand sorting, gravity separation, floatation, magnetic separation, acid leaching, etc. Usually barite associated with oxide ore such as fluorite dolomite, quartz ore etc. are also associated with sulfide ore such as pyrite, sphalerite and galena. The separation between barite and sulfide ore (such as pyrite) is easy. On the other hand, the separation between barite and oxide ore (such as dolomite and fluorite) is very difficult. There are quite a lot of reports on the separation between barite with fluorite dolomite however the separation between barite and dolomite is seldom discussed.The ore properties are quite complex attributable to mineral assemblage containing a large amounts of dolomite. As the density of barite and dolomite has some differences, gravity separation between them is usually easy. The gravity separation tests were conducted with no results obtained due to the fine dissemination size in a complex sedimentary. The magnetic separation was conducted to remove the magnetic gangues, and the results are not satisfactory as the gangue minerals such as limonite and others were low in content. Due to the fine dissemination and intimate association of minerals, and barite and dolomite contain the same group ions of Ba2+, Ca2+ and Mg2+, the results of flotation tests showed that the mineral separation between barite and dolomite by flotation means problematic. Previous studies have been conducted by several institutions but none to date has come up with a satisfactory process in dealing with the deposit. This may be attributable to: high cost of reagents; unsatisfactory performance. Therefore, the key to the flotation separation of barite and dolomite is the application of selective inhibitors. Based on previous test observations and years of extensive research experience in barite processing technology, with the new depressant TS has selectivity inhibition on dolomite developed and used, good results can be obtained by “desulfurization-barite flotation”, which allows barite in the lead-zinc tailings to a comprehensive recovery."

Jan 1, 2018

By Vasily V. Lodeischikov, Adolf F. Panchenko, Olga D. Khmelnitskaya, Vladimir E. Dementyev

Pilot plant tests of underground gold leaching from Maminskoye deposit oxide ores at a grade of 0,7 g/t Au were carried out. The underground leaching flow sheet includes Au dissolution by chlorine-containing solutions, Au adsorption from solution on active carbon, carbon concentrate ashing and calcine melting to produce high grade gold concentrate. The underground leaching system throughput was 15 m3 of solution per hour. Gold concentration in pregnant solutions was 0.12-0.13 mg/l. Gold recovery from solutions into carbon was in the range of 63-84%. It was demonstrated that the low Au recovery in the adsorption cycle was due to the presence in pregnant solutions of "active" chlorine, which destroys active carbon and causes a sharp decrease of the carbon loading capacity for gold. To improve the adsorption indices, previously dechlorinated pregnant solutions using bituminous coal or Na2S2O3, is recommended. Ion exchange process was tested as an alternative for gold extraction from dechlorinated solutions. In the testwork period, 4.4 kg of gold were produced with a chlorine consumption of 3.5 kg per 1 g of gold. During the tests a program of environmental protection measures was completed to provide the necessary safety precautions of the underground leaching works.

Jan 1, 2003

By Y. Y. Zhou

This paper introduces the application of CRIMM high-gradient magnetic separator in refinement of feldspar powders as well as the supporting technical outfits needed by the refining process with the high-gradient magnetic separator as main equipment. It could reduce the content of Fe2O3 to 0.03% with a deferrization rate of 85.88% from the feldspar raw ore containing 0.17% Fe2O3by using the above refining technology, producing an ultrapure feldspar product.

Jan 1, 2014

By Ernesto de la Torre, Karla Velín, Alicia Guevara

"The following study shows the improvement and simulation of a gold recovery plant of 24 t/day of capacity by gravity concentration, flotation and leaching. The ore processed had grades of 6 g/t Au, 15 g/t Ag, and 0.69 %Cu with the mineralogy consisting of 2 %chalcopyrite, 12 %pyrrhotite, 7 %plagioclase and 50 %quartz. The recovery process starts with crushing the ore followed by grinding it further in a ball mill. The ground ore is then concentrated in a Falcon centrifugal concentrator with concentrate being processed by amalgamation process. Gravity tails are sent to a flotation circuit where the copper minerals and some gold are concentrated. Lastly, the flotation tailings undergo cyanidation for gold recovery. Currently, the overall recoveries are 82.8 %Au, 49.6 %Ag and 99.1 % Cu. The optimization was developed in the laboratory using a Wilfley table by gravity concentration. Flotation tests in the laboratory were performed using potassium amyl xanthate and pine oil. Lastly, cyanidation was then performed on the flotation tailings. The results showed the optimum concentration of sodium cyanide in cyanidation is 2 g/L, which produced an increase in gold recovery from 30% to 44%. The reduction of diameter particle from 132 to 80 µm also produces an incremental recovery of gold and silver from 11% to 21% Au and 0.4 to 3% Ag in gravity concentration. Likewise in the cyanidation the recovery of gold and silver has shown a rise from 44 to 62% and 8 to 32% for Au and Ag respectively. However, the flotation method the gold recovery is reduced from 68 to 30% Au by reducing the particle size. Computer simulations using the programs, Moly-Cop and ModSim were performed to determine optimum parameters for the recovery of precious metals at the gold recovery plant."

Jan 1, 2014