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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 Agnieszka Surowiak, Tomasz Niedoba

The paper presents a way of determining relations between ash content in coal, particle size and particle density by means of regression, multidimensional Morgenstern distribution functions and kernel approximation. The three-dimensional random vector (D, ?, ?), where D is a random variable describing particle size, ? - random variable describing particle density and ? - random variable describing ash content, where function ?=?(d, ?) is obtained by means of regressive function. The multidimensional Morgenstern distribution functions and kernel estimators with the use of Gauss function were applied to approximate the functions f(d, ?, ?) and f1(d, ?). This approach is a new way of multidimensional description of grained materials such as coal and it also allows to evaluate material features more properly than the classical statistical approach. The results were verified statistically.

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 I. N. Gridasov, V. V. Khromov, R. I. Ismagilov

"Over 50 mln.tonne is the average annual output of oxidized ferruginous quartzites typically containing Fe38-42% mined in Mikhailovskoye ore field. The most promising and cost-effective way to expand concentrate production maintaining current output of magnetite is processing of oxidized ferruginous quartzites (hereinafter referred to as OFQ) mined as a by-product in Mikhailovsky open pit mine. However lack of proven and efficient concentration and agglomeration process applicable to OFQ and insufficient data for feasibility evaluation of OFQ processing encumbered full-scale addressing of this challenge. Based on outcomes of multiple various studies a process of OFQ flotation was designed by MC Metalloinvest LLC and PJSC Mikhailovsky GOK engineers. Comparative lab-scale testwork have proven performance and production advantages of flotation against magnetic separation feat. flotation and combined concentration method. Lab-scale flotation enabled production of concentrate with average Fe66.6% and SiO22.06% grade. Geologic and production contouring of OFQ deposited in Mikhailovskoye ore field has been carried out to collect necessary data for feasibility study incl. cost and profit estimation. As a result processing and production properties of over 3,500 samples have been analyzed which is characteristic of every direction and level of prospective ore mining within Mikhailovskoye ore field up to year 2032. Following work has been done based on geologic and production contouring of OFQ: – natural typification of ores and production (processing) grades of OFQ, determination of average mineral composition; – correlation analysis of interrelation between OFQ mineral composition and processing features thereof for overall bulk of samples as well as for individual data blocks related to various directions of open pit mine; – OFQ classification with account of defined mineral composition and processing features. Total Fe weight content in chamber product of flotation (known as concentrating ability in flotation) was applied as a major differentiating factor for processing grade classification of OFQ. Three processing grades of oxidized ferruginous quartzites were established as a result of flotation tests run to determine concentrating ability, i.e. – hard concentrated OFQ (TO) – below Fe65%; – mean concentrated OFQ (SO) – Fe65-67%; – easy concentrated OFQ (LO) – over Fe67%. Outcome of this work provided initial data for feasibility study and cost-profit evaluation of medium-term industrial-scale processing of OFQ deposited in Mikhailovskoye ore field."

Jan 1, 2018

By A. M. Raichur, S. P. Vijayalakshmi

Fine coal dewatering is a major problem in the coal industry because the fines hold significant amounts of water and account for two-thirds of the product surface moisture. Present methods include flocculation using polymeric flocculants followed by filtration. But the kinetics of filtration are not favorable with such flocculants. In this study, a novel method involving flocculation with microorganisms for improving the settling rates and dewatering kinetics has been investigated. Two Gram-positive microorganisms, Mycobacterium phlei and Bacillus subtilis were used as bioflocculants. The effect of time, pH and ionic strength of the electrolyte used on adhesion and flocculation were studied. Adhesion tests showed that maximum adsorption occurred in 20 minutes in the pH range of 2-7 and the total amount of bacteria adsorbed was almost the same with both the flocculants. Nearly 98% of the coal flocculated with both the microorganisms in about 5 minutes. The filtration rate of fine coal almost doubled with addition of bacteria and electrolyte and maximum filtration occurred near the point-of-zero-charge of the bacteria.

Jan 1, 2003

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 Hao Zhang, Dezhou Wei, Qiang Zhao, Caie Zhang, Baoyu Cui, Xuetao Wang

"Nowadays, long periods of relatively low iron ore prices have made the steel industry get the greater pressure. To develop appropriate processing technology based on characteristics of the iron ore, so as to reduce the beneficiation costs, is crucial to the improvement of iron and steel enterprise competition. For a typical low-grade iron ore from northeast China, beneficiation process research was conducted based on the in-depth studies of ore characteristics. Results show that the iron grade of the ore was 31.56%, the content of FeO was 6.59%, the main iron minerals were hematite and magnetite with fine dissemination sizes and partly scattering in gangue; With the grinding fineness of 65% passing 0.074mm, and regrinding fineness of 90% passing 0.074mm, the concentrate with 66.80% iron and recovery of 82.90% was obtained via the combined flowsheet of stage grinding – classification – gravity concentration – low and high intensity magnetic separation – reverse flotation. Ratios of gravity concentration and low intensity magnetic separation were 70.21% and 7.57% respectively in 40.42% of total concentrate yield; The loss of micro-size fractions can be reduced by feeding spiral tailings into magnetic separation and flotation; It’s helpful in reducing the dosage of reagents and beneficiation costs by increasing the throughput of gravity concentration and outputting concentrate of low intensity magnetic separation.INTRODUCTION Iron and steel industries are facing the increasing pressures in the continued falling iron ore prices. In order to improve the competitiveness, it is extremely urgent for them to develop more appropriate separation process for the reduction in iron ore beneficiation cost. As a typical low-grade iron ore, Anshan-type iron ore is one of the most important iron ore resources in China [1-3]. The flowsheet of stage grinding and separation contained the classification of the coarse particles and the fine particles, the gravity concentration, the magnetic separation and the anionic reverse flotation has been widely applied for Anshan-type iron ore. With the appropriate separation progress as well as equipment, satisfactory separation results had been obtained, and mineral particles with different size were recovered separately [4-7]. However, the low recovery of fine mineral particles and the relatively high beneficiation costs still prevented the efficient application of the process [8]. For a Anshan-type low-grade iron ore, a typical process flowsheet, as mentioned earlier, with stage grading, classification of the coarse particles and the fine particles, gravity concentration, magnetic separation, anionic reverse flotation was adopted currently. And the concentrate with 67.10% iron and recovery of 75.50% was obtained and the gravity concentrate accounted for 39% of the total concentrate. However, a large amount of fine liberated iron ore particles were lost in the tailings."

Jan 1, 2018

By G. Demi, G. Belardi, N. Shehu, S. Cico

The physico-chemical interaction between minerals such as olivine, chromite and serpentine and the ions most commonly present in solution, like Ca and Mg and flotation reagents such as EDTA and oleic acid was investigated. The study was conducted via release tests, electrophoretic measurements and microflotation of selected minerals coming from the Kukes deposit (Albania). The results were confirmed by means of beneficiation tests on various types of ores, from the Kukes (Kalimash), from Bulquiza area and from the Rizo deposit (Greece). It has been demonstrated that when Na-K Oleate is used as collector chromite, olivine and serpentine can be separated at pH = 2.5 and/or 5.5 employing EDTA (EthyleneDiamineTetraAcetate) on previously deslimed material. This approach ensures flotation concentrates with a Cr2O3 content about 50%, at a recovery in excess of 65%.

Jan 1, 2003

By I. I. Maximov

The paper describes the results of studies performed with a view to develop an improved technology for processing high-grade massive and cupriferous ores of the Talnakh deposit, the main valuable components being nickel, copper and PGM. The studies? main task was production of nickel concentrate with nickel content of approx. 15%, nickel recovery being as high as possible. The studies were performed on two ore samples. One sample was a mix containing high-grade massive (78%) and cupriferous (22%) ores with nickel content of 1.93% and copper content of 3.6%; the other sample was a mix of high-grade ores with nickel content of 2.25% and copper content of 3.72%. Main ore minerals were copper sulfides ? chalcopyrite and cubanite, their total content being 10.5% ~ 11%, pentlandite (5.2% ~ 6.1%) and pyrrhotite (36.7% ~ 42.9%). Influence of grinding media on pentlandite and pyrrhotite separation results, achieved through combined use of triethylenetetramine (TETA) and sodium sulfite (SS), was also investigated. It is shown, that combined application of TETA/SS in the ratio of 1:2, permits to grind products in metal mill (mild steel) with metal balls (mild steel), achieving results that are similar to those achieved by means of grinding in inert grinding media with TETA/SS ratio of 1:1. Application of the developed technology provided for production of high-grade nickel concentrate with nickel content of 14.7% ~ 15.6%, nickel recovery being 76 % ~ 77%. Earlier experiments performed on the above mentioned ore mix permitted to produce nickel concentrate with nickel content of 10%, nickel recovery being 75.9%. Consequently, nickel content was increased by 4.7% ~ 5.6%, and a higher nickel recovery was achieved. The paper also describes metallurgical results related to nickel, copper and PGM with regard to all processing products.

Jan 1, 2014

By A. E. Sobolev

The kinetics of hydrochemical oxidation of antimonite in solutions of nitric acid was studied using the rotating disk method and the factor experiment design. The dependences of the antimonite?s oxidation rate on concentration of nitric acid solutions, temperature, disk rotation frequency, and duration of interaction were determined. The adequate kinetic models allowing calculating the specific oxidation rate at any combination of the values of the influencing factors were obtained. The oxidation process is found to proceed in a mixed mode, close to chemically controlled. Gradually, as the sediments of elemental sulphur and Sb2O5 accumulate on the disk surface, the process becomes controlled by internal diffusion. The observed high (more than second) order of the oxidation rate with respect to the concentration of nitric acid was explained by the autocatalytic action of the acid reduction products. The apparent activation energy depends on the stage of interaction. At the initial stage (in a mixed mode of the process proceeding), it depends on nitric acid concentration and varies from 31.5 kJ?mol?1 (in 2 M HNO3) to 36.1 kJ?mol?1 (in 10 M HNO3). This trend indicates the weakening of diffusion limitations with a nitric acid concentration increase. When the process is controlled by internal diffusion, the apparent activation energy was found to be 14.5 kJ?mol?1. The obtained kinetic characteristics enlarge the database on hydrochemical oxidation kinetics of sulphides and can be used for the elaboration of the recommendations for the effective comprehensive exploitation of antimony-bearing ores and concentrates.

Jan 1, 2014

By A. Mainza, H. Hill, L. Bbosa

"The main objective of ore breakage characterization is to quantify the relationship between the energy supplied for breakage and the size of the resultant progeny. In this study, the ore breakage characteristics of drill core particles and crushed particles were compared. Single impact breakage tests were conducted on a basalt ore sample using the JK rotary breakage tester. The size-dependent model by Shi and Kojovic was used to fit the breakage data because the effect of initial particle size is accounted for in the model. It was found that crushed particles consistently gave a higher degree of brokenness than drill core particles and the differences were attributed to the internal structure of the particle. Drill core particles contain a lower crack density (paths of weakness) than crushed particles which have been pre-weakened during the crushing process. An increase in crack density is directly related to a decrease in the minimum threshold energy. Upon application of a stressing force above the stress threshold, particles with a higher crack density will fracture more easily because of a higher rate of crack propagation due the numerous paths of weakness available. It was therefore concluded that under the same conditions, drill core particles are more competent than crushed particles. INTRODUCTION Comminution devices have commonly been reported as being the largest energy consumers in mineral processing circuits (Napier-Munn, 2015). Cleary (1998) reported that the inefficiency of comminution devices is a result of less than 5% of the supplied energy being utilized for breakage, with the majority being expended as heat and noise. With steadily declining fossil fuel supplies and a looming global energy crisis, the mineral processing industry has increased its focus on energy reduction measures. Ore breakage characterization plays a critical role in the design and optimization of comminution circuits (Tavares and King, 1998). The main objective of ore breakage characterization is to empirically quantify the relationship between the energy supplied for breakage (specific input energy) and the size of the resultant progeny through laboratory particle breakage tests (Napier-Munn, 1996). Commonly used laboratory breakage devices include the Drop Weight Tester (DWT), Split Hopkinson Pressure Bar (SHPB) and more recently, the JK Rotary Breakage Tester (JKRBT). Typically, a standard single impact breakage test procedure consists of a matrix of experiments on particles in four to five narrow size fractions each tested at low, medium and high energy levels. The resultant progeny of each test is sized to obtain a particle size distribution (PSD) from which the degree of breakage, usually in the form of a t10 value (percentage of the progeny material that passes 1/10 of the feed size) can be obtained for each size fraction and energy level (Napier-Munn, 1996)."

Jan 1, 2018

By Wei Sun, Ruolin Wang, Jianjun Wang, Xiaodong Li, Honghu Tang, Yuehua Hu, Anh V. Nguyen, Haisheng Han, Kefeng Chen, Zhao Wei

It is difficult to separate scheelite especially from fluorite with fatty acid collectors because the adsorption mechanism of these collectors occurs through chemisorption of the oleate ion onto the mineral surface. A simplified flotation process using Pb-BHA complexes was developed to replace the energy-inefficient Petrov’s process of using a fatty acid, increasing the recovery of scheelite and reducing the cost. The relationship between surface chemical properties (crystal chemistry, solubility, and Zeta potential) of minerals and flotation by using a fatty acid or Pb-BHA complexes was studied in detail to provide new insights into the selective separation mechanism. The results indicate that the scheelite surface exposes more O atoms than Ca and that they are negatively charged due to the anisometric dissolution of Ca and WO4, which provides the possibility for separation using selective interfacial reaction. The FTIR and Zeta potential results indicate that both chemical and electrostatic adsorption play an important role in the adsorption of Pb-BHA complexes on the scheelite surface. The XPS results of the atomic composition and binding energies confirmed the presence of the adsorbed Pb-BHA complexes and that O and Ca may be the active adsorption sites. Therefore, scheelite was able to be collected by forming either a Ca-O-Pbcollector or W-O-Pb-collector (BHA). However, fluorite was only able to be collected by a fatty acid collector and colloidal [H2SiO3]n -, and a high temperature was needed improve the selective desorption of the adsorbed fatty acid collector by forming Ca-[H2SiO3]n - on the fluorite surface, resulting in serious problems in industry. However, the Pb-BHA complexes significantly contribute to improving the recovery of scheelite minerals, simplifying the flotation process, and circulation of the residual reagents and water. The new process has been applied in Shizhuyuan Mine, and improved the recovery by almost 10%.

Jan 1, 2018

By Klaus M. Hahn, Kurt Aasly, Jutta Lennartz, Hermann Wotruba, Ben Snook, Przemyslaw B. Kowalczuk, Rolf Arne Kleiv

"Seafloor massive sulphide (SMS) deposits have been identified as important marine metal resources for the future. However, literature on mineral processing of SMS is currently limited, and to date, no research has been published on processing of SMS from either active or inactive hydrothermal vent fields on the Arctic Mid-Ocean Ridge. In this paper, individual SMS rock samples from the Loki’s Castle area were tested. Chalcopyrite, isocubanite and sphalerite represented the main copper and zinc bearing minerals in the SMS samples, whereas the gangue consisted of barite and silica. Based on the X-ray transmission response (XRT), i.e. differences in the atomic number of elements in tested samples, and using portable XRF (PXRF) values of elemental composition, individual particles were classified as a either product (low barium / high to low silicon / high copper and zinc) or waste (high barium / high silicon / low copper and zinc). The classification was based on the elemental composition of the samples with the cut-off grade equal to 0.5% for copper. It was shown that the XRT discrimination analysis correlated well with elemental (PXRF) data classification. The results clearly indicate that X-ray transmission sensor-based sorting can be used a useful technique for preconcentration of SMS rock samples. INTRODUCTION Due to different compositions of investigated samples, sensor-based sorting is a useful technique for preconcentration of different materials. The aim of using sensor-based sorting is to upgrade an ore by reducing the quantity of overall waste material reporting to the next processing stage. This can be achieved by eliminating (reducing) the waste fraction. By applying sensor-based sorting, the energy consumption for particle comminution as well as the required process water and reagents can be decreased. The total volume of the waste rejection leads to a proportional reduction in the total energy costs (Lessard et al., 2014, 2016)."

Jan 1, 2018

By Milind Chikhale, Ranjeev Kumar, Keshav Tyagi

Ore characterisation including mineralogy and beneficiation testwork of an ore is very important for determination of possible products and their quality. The results of such study help in ascertaining the beneficiation process route for the ore. The present study deals with ore characterisation study and beneficiation testwork on the samples of hematite iron ore with goethite impurity. The characterisation study reveals that the hematite is under alteration stage to goethite which has become an integral part of hematite. High goethite/limonite content is enhancing the loss of ignition content of the ore and reducing its Fe content. The findings indicate that in case of hematite iron ore associated with goethite and limonite, it is difficult to achieve high grade of concentrate above 65% Fe. It depends on the amount of goethite and limonite minerals and their association with hematite. These minerals contain aluminous impurities and hence, their rejection becomes essential. Separation of limonite and of goethite to limited extent is possible. In case of goethite, effective liberation is not achieved where it occurs as surface alteration product of hematite grains. The gangue mineral reported are clay and gibbsite, and are closely associated with iron minerals. The test work results indicate the grade and recovery of Fe along with alumina and silica content in lumps, sinter fines and pellet feed fines that can be achieved for the different ore types vis-a-vis the percent of available hematite in the ore. The recovery of iron is dependent on the hematite available and the loss of iron values in tailings is more because of rejection of limonite and goethite.

Jan 1, 2014

By Yimin Zhu, Dezhou Wei, Haijun Liang, Zhifu Zhou

It has been fairly well established by many experiments that Thiobacillus ferrooxidans (T. ferrooxidans), also named Acidithiobacillus ferrooxidans(Kelly and Wood, 2000),can depress the floatability of pyrite. However, the mechanisms governing the process have not been known completely. In the present investigation, wettability and electrokinetic studies were carried out to examine the surface chemical changes on pyrite before and after interaction with T. ferrooxidans or its secretions. Attachment of T. ferrooxidans to pyrite was observed using a scanning electron microscope. The results of experiments showed that, under certain conditions, T. ferrooxidans and its secretions would obviously change the surface electricity of pyrite and result in the iso-electric point of pyrite shifting notably. However, the surface wettability of pyrite hardly changed and its hydrophobicity increased little after interaction with T. ferrooxidans and its secretions. Based on the results of the present investigation it can be reasoned that the depression of floatability of pyrite is due to the reduction of surface free energy, rather than the increase of surface hydrophilicity of pyrite. The decrease of surface free energy weakens the attachment of collector to pyrite; during the process bacterial protein plays the most important role and the bacterial chemotaxis strengthens the effect.

Jan 1, 2003

By S. V. Balikov

A close association of fine gold with pyrite and arsenopyrite is observed at many gold ore deposits. Gold forms and gold distribution in sulfides of different deposits were studied. A comprehensive method including optical spectrum analysis, X-ray diffraction analysis, high-temperature X-ray diffraction analysis, scanning electron microscopy and X-ray microanalysis was used for the de-tailed investigation of the defectiveness of the structure of pyrites from different deposits. It was shown that gold content in pyrites varied from 1.5 g/t to 100 g/t compared with 30 g/t to 230 g/t in arsenopyrites. The nature and arrangement of inner inclusions of foreign minerals in pyrites as well as occurrence of fractures and macro- and microdefects were examined. Iron, sulfur and arsenic contents in all pyrite samples were determined. Chemical composition of the detected inclusions was defined. Parameters of submicroscopic structure of pyrite were estimated using X-ray Fourier line shape analysis. Estimation of real structure characteristics with calculation of submicroscopic D-blocks and root-mean-square lattice deformation 2()evalues was performed. Defectiveness of the pyrite structure can be defined by the following parameters: K?1 and K?2 reflection resolution and its intensity within high Bragg angles; S content to Fe content ratio; the arsenic content and the local variation in its concentration within the pyrite grain; the size of submicroscopic blocks and the crystal lattice disturbance value. The performed analysis showed some relationships between these parameters and refractoriness of the pyrite concentrates. It was observed that at S/Fe ratio close to or higher than stoichiometry and at a low arsenic content (less than one per cent) pyrite concentrates were not refractory. The analysis results showed that concentrates of the Beryozovskaya gold plant (0.02~0.07 As%) and concentrates of the Kommunarovskaya gold plant (less than 0.05 As%) as well as some pyrites from the Sukhoi Log gold deposit had the same characteristics. It was found that average arsenic content did not excess 0.2 per cent. Arsenic distribution within the grain of all these pyrites was uniform. At S/Fe ratio lower than or close to stoichiometry and at high arsenic content (more than one per cent) and significant variation in its distribution within a pyrite grain the pyrite concentrates are refractory. Such pyrites are typical for the Sukhoi Log, the Bakyrchik and the Mardzhanbulak gold deposits. Direct correlation between arsenic and gold contents, in particular, in fine grains of pyrite was observed by many researches. This correlation was explained by impregnation of arsenopyrite which normally contains more gold than arsenic pyrites. The analyses results showed that no arsenopyrite impregnation was found: arsenic distribution within the grain was uniform but its concentration varied greatly. Such pyrite concentrates are more refractory. This paper presents the results of analyses of sulfides to determine the amenability of the associated gold recovery by hydro- and pyrometallurgy.

Jan 1, 2014

By Andrzej Wieniewski, Cezary Regula, Bozena Skorupska, Krystian Cichy

"In this paper, experiences from Polish non-ferrous metals ores enrichment plants in the copper and blende concentrates quality improvement are discussed. The blende concentrate being produced in ZGH Boleslaw S.A., with Zn content of about 54,5% is characterized by the relatively high magnesium content reaching a level of 1% resulting from the dolomite presence, while for the zinc production by electrowinning a concentrate in which magnesium content should not exceed 0,3% is required. The implemented new technology for the quality improvement of blende concentrate is based on the leaching of zinc concentrate with sulfuric acid aimed at dissolving of the contained dolomite, and then flotation of the leached concentrate. It allows to obtain a concentrate with magnesium content below 0,1% at the simultaneously increased zinc content to about 57,5%. In the paper, laboratory tests results on the development of optimal flotation flowsheet for the gypsum removal and the industrial tests results concerning the process optimization of this flotation are discussed. In turn, the copper concentrate being produced in Polish copper ore enrichment plants KGHM PM S.A., is characterized by the increased organic carbon content - above 8%. It creates problems with the smelting of this concentrate because of difficulties with the heat removal derived from the combustion of organic carbon in the flash smelting process. The implementation of segregating flotation to the industrial practice allows the effective production of two concentrates with the differentiated chemical compositions - a concentrate with increased copper content and reduced organic carbon content for the flash smelting process, and a concentrate with lower copper content and increased organic carbon content for the shaft smelting process. In the paper, conditions of the segregating flotation and the achieved results are described."

Jan 1, 2014

By A. V. Bogdanovich

In diamond ores processing, final waste tailings size, rejected to dump, is one of the most important parameters. On one side, it is necessary to minimize losses of diamonds, whose size exceeds 1 mm, on the other side ? to provide for the plant high throughput and sufficiently low capital and energy costs, which is in greater part connected with disintegration process. Extensive laboratory, pilot-scale and commercial-scale dressability studies of different types of diamond ores have been performed at JSC ?Mekhanobr Engineering? in recent years. With that, special consideration was given to determination of the required ores? disintegration size that would provide for a high recovery of diamonds. In diamond ores processing, in order to exclude breaking of large diamonds, stage disintegration is used with regrinding of light fractions and heavy fraction X-ray fluorescent separation tailings. On account of rather weak bonds of diamonds with enclosing minerals, in crushing and grinding of diamond ores, selective disintegration takes place, with aggregates? breakup along plane surfaces of least bonding, which facilitates liberation of diamonds and their transition to free state. Heavy components losses with different sizes of final dump tailings were studied in processing of large samples. It was established, that in order to almost completely exclude losses of heavy components over 1 mm size, light fractions had to be ground to 2.5-3 mm size, while X-ray fluorescent separation heavy tailings of -15+3 mm size do not contain diamonds larger 1 mm in aggregates, and their regrinding does not increase total recovery of diamonds. On the basis of the conducted studies, relationships regarding heavy components and diamonds losses with respect to diamond ores final disintegration size were obtained, providing for more rational organization of the ores disintegration and processing flow sheet. Keywords: diamond, waste tailings size, disintegration process, recovery of diamonds

Sep 1, 2012

By Bruce Mathema, Steven Huynh, Terry Cheng, Gerry Gomez

"The use of PM40N, a specialty precious metals promoter in conjunction with a standard Xanthate based collector was extensively investigated in the flotation of a complex, silver-bearing, lead-zinc ore containing naturally occurring graphitic gangue. The ore, considered difficult to treat with conventional reagents, was tested at plant scale to determine the impact of the promoter on silver recovery, lead grade and also on the amount of zinc misplaced to the final lead concentrate.Initial laboratory-scale rougher flotation testing implied significant improvements in the recovery of silver and lead with a marginal amount of zinc displaced. Subsequent mini pilot flotation tests further revealed that a gain in selectivity and recovery could be achieved by adjusting PM40N promoter dosage and incorporating a cleaner flotation step, further improving rejection of sphalerite. Plant trials resulted in an improvement of 3% in Silver recovery coupled with a 2% increase in recovered lead whilst maintaining lead grade in the lead circuit.To date, the use of PM40N as a promoter has enabled the processing operation to significantly increase its plant economics. The partial replacement of the standard Xanthate based collector has also reduced the overall quantity of Xanthate required in the standard circuit by approximately 10% - which is a positive impact on the environment. It is envisaged that further optimization of plant parameters will favor stable operation and continue to boost revenues generated by subsequent improvements in silver metallurgy."

Jan 1, 2014

By Vitaly Kretov, Valery Golikov, Vladimir Ryaboy, Elena Shepeta

"The aim of the study was to develop and implement more efficient dialkyldithiophosphate collectors. A hypothesis has been formulated on the action of dialkyldithiophosphates in the flotation of sulphide ores based on the assessment of their complexing properties and surface activity at the liquid-gas border. Dialkyldithiophosphates form weaker complexes with cations of heavy metals compared to xanthates, and are characterized by a large difference in the strength of their complexes with heavy metal cations. It provides higher selectivity of dialkyldithiophosphates in the separation of non-ferrous metal sulphides and iron sulphides from them. To make a quality assessment of the impact of surface activity of the synthesized collectors on their selectivity of action and collecting power, an assessment of the surface tension of their aqueous solutions has been performed using the stalagmometer method. It is well known that a stronger collector is characterized by a greater surface activity, in contrast to a more selective collector which is characterized by a lower surface activity. Based on studies, new dialkyldithiophosphate collectors BTF 163 and BTF 1552 are suggested for use in the flotation of sulphides to increase the recovery of copper, gold and silver. The collector BTF 163 was introduced at a gold processing plant in Kazakhstan and has been undergoing industrial testing in the sulphide cycle of scheelite-skarn-sulphide ores in a factory in the far east of Russia. The collector BTF 1552, which increases the recovery of silver and gold by 3%, was approved for industrial testing in another far-eastern factory. A modification of the reagent BTF 1552, the new reagent BTF 1522, was introduced in the flotation of copper-zinc ore at a factory in Kazakhstan."

Jan 1, 2014