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By K. D’Elia, L. D. Ogrel, O. A. Malyshev, T. I. Yushina, I. M. Petrov

"The purpose of this research is to develop enhancements in the mineral processing by flotation methods of nonferrous ores. Currently, 180-190 million tons of ore are processed by flotation in Russia. The annual processing volume of copper-nickel and copper-zinc ores exceeds 20 million tons. In the total volume of processed ores by flotation, non-ferrous ores account for roughly 35% and ores of precious metals for another 15%. The total demand of flotation reagents in Russia in 2016 was estimated at 170,000 tons. Sulfur-based collectors used in the flotation of non-ferrous sulfide ores (typically, xanthates) do not always offer sufficiently high recovery of valuable components.It has been found that reagents like DMIPEC and DK-80 improve the recovery of metals into concentrates. To identify the reasons of that higher recovery, quantum chemical calculations of the mechanism and interaction forces of DMIPEC with non-ferrous sulfides were undertaken by determining the stabilization energy of the resulting chemical complexes. It has been shown that, in the interaction of DMIPEC and DK-80 with the surfaces of sulfide minerals, ethylene and acetylene bonds act as the reaction centers, and the formation of molecular p-complexes is the result of metal-cation interactions with the surface and the flotation reagents.In the flotation of non-ferrous ores, DMIPEC and DK-80 act as selective supplementary collectors with foaming properties, and at a flow rate of 3-35 g/ton, they reduce the flotation time, and increase copper recovery by up to 3.5%, zinc recovery by up to 12%, gold recovery by up to 8-10%, allowing to increase the recovery of valuable components by 1.5-12.0%, basically this kind of reagents nearly double the flotation rate, improve the grade of the concentrate, improve the stability of the flotation process parameters, and reduce the environmental impact of flotation plants."

Jan 1, 2018

By R. P. Bhagat

The paper reports results on sinterability of iron ores from the mines of Tata Steel Limited. The objective was to improve tumbler index (TI) of sinter and its reduction parameters, reduction degradation index (RDI) and reducibility (RI) while maintaining strand productivity same or to improve. Pot sintering experiments were carried out with Joda and Noamundi mines having LOI 3.8% and 2.9% respectively. Sinter basicity (CaO/SiO2) was varied in the 2.1 to 2.5 range and its MgO content was kept in the 1.4% to 1.8% range according to the blast furnace requirement. The ignition condition and vacuum was maintained according to the plant practice. The effect of variables namely, granulometric composition of ores vis-à-vis that of sinter mix, bed height, solid fuel, moisture content of green mix on the sintering indices was studied. The productivity with Noamundi ore (26 %+6.3 mm) was lower than that with Joda ore (45% +6.3 mm) at 4% coke in the mix and sinter basicity (CaO/SiO2) 2.1. However, with 4% coke (ash 22%) the RDI value was significantly high in case of Noamundi Ore, which decreased significantly to 29% with 5% coke. The size parameter of ore fines and its LOI content do influence the sintering speed and the sinter properties significantly. At 2.1 basicity ratio, the sinter with the Joda ore having coarsergranulometry had highest TI (=70.2%). The RDI (27%) in this case was inferior to that (23.2%) when Joda ores with finer size was sintered. Better RDI with finer sized ore could be attributed to better assimilation of CaO/ MgO with more balanced and uniformly distributed hematite minerals. The size parameter has been inter-influenced by the coke requirement. With the lower mean sized Joda ore the sinter RDI decreased significantly to 23% and the productivity improved to 42.9 T/(sq m*d) when the coke in the mix increased from 4% to 5%. An attempt to reduce the coke in the mix from 5% to 4% with respect to both the ores was not successful as it resulted higher RDI and lower strand productivity. The present study has revealed that sinter RDI improves (decreases) with decrease in the sintering time, while such correlation was absent between TI of sinter and speed of sintering. Corollary to this observation, one can take advantage of other stronger parameters to increase the sintering speed (and hence sinter productivity) without affecting the sinter strength. In the present study higher FeO sinters has displayed excellent RDI, however cold strength of sinter was not correlated with its FeO content. The TI and RI parameters of sinter improves marginally with increase in its MgO content while RDI got worsened. Keywords: LOI content, iron ore, reduction properties, sinter, noamundi, joda

Sep 1, 2012

By E. V. Zelinskaya, V. U. Starostina

"The Irkutsk region is a raw material region. There are a lot of big enterprises for extraction and processing of oil, gas, coal, iron, gold and rock salt. Besides of it, refractory clays, limestone, mica, gypsum and talc are mined in the region. The most important type of mineral raw material is a gold. More than 8.5% of the total mining in the Russian Federation is extracted in Irkutsk region. It is about 0.7% of the world production. The environmental problems of mining and processing enterprises are a complex set of issues that requires of the searching of effective methods to be solved. One of the key environmental problems remains the reduction of waste generation and the introduction of environmentally safe waste management. After analysing different methods of environmental impact assessment it is established that the method of Life Cycle Assessment is one of the most effective methods for mining production. This method base? on the material balance of flows of raw materials, products and waste. The method of Life Cycle Assessment allows to identify the effects of environmental pollution under the influence of mining and processing at a quantitative level and to show real quantitative forecasts of the possibility of managing the flows of raw materials and waste.INTRODUCTION The Irkutsk region is a raw material region. There are a lot of big enterprises for extraction and processing of oil, gas, coal, iron, gold and rock salt. Besides of it, refractory clays, limestone, mica, gypsum and talc are mined in the region. The most important type of mineral raw material is a gold. More than 8.5% of the total mining in the Russian Federation is extracted in Irkutsk region. It is about 0.7% of the world production. The main provider of precious metal is the Bodaibo region. This is a main location for the major ore companies such as “Vysochaishy”, which produces about 5.0 tons a gold per year; “Pervenets”(PJSC Polyus, the largest gold producer in Russia) - 4.5 tons/year, “Druza” - 0, 5 tons/year. Placer gold is extracted by more than 40 enterprises. In connection with this, there are many plans about a building of new and reconstruction of concentrating factories. Consequently, an increase in the volume of processing is expected. The main projects are the construction of a new mill of the Mining and processing plant GOK “Vysochaishy” with a capacity of 2.6 million tonnes of ore per year; the development of the “Uhakhan” gold-field - 23 km from the “Goltsa Vysochaishy” gold-field; the opening in the Bodaibo region the next parts of the gold-deposit ""Pervenets"". Additional, the growth in a gold production can be achieved by a store of gold in the “Sukhoi Log” deposi"

Jan 1, 2018

By Sanjeeva Latchireddi

The effective utilization of the energy in comminution devices depend on how efficiently the broken particles leave the breakage field as soon as they reach product size. In crushers the transport of broken particles smaller than CSS is done by gravity (vertical) thus allowing the energy to be applied to new particles. However, in AG/SAG mills, gravity becomes the energy field and the transport of particles is horizontal which needs a suitable discharge mechanism and has been found to have significant effect on grinding efficiency. The material transport in AG/SAG mills depend on how efficiently the discharge system (grate and pulp lifters) operates. Material transport plays a significant role on efficiency of horizontal grinding mills ? AG/SAG/Ball mills. The magnitude of benefits that can be achieved with efficient material transport in AG/SAG mills has been demonstrated in a Greenfield milling operation at Mirabela Mineração?s SantaRita Nickel operation located in northeastern part of Bahia in Brazil. This grinding circuit was commissioned in October 2009, which was designed as a typical SABC circuit to grind 4.6 mtpa (575tph) to produce a P80 of 125um with 10-12% ball charge in SAG mill. However, it has been operating as ABC circuit without any grinding media in SAG mill. The nameplate design capacity has been achieved in few days and has gradually increased to an average of +800 tph and still it is operating in ABC mode. The milling circuit commissioning resulted in small and relatively simple measures to rectify, which was reflected in the rapid commissioning of the milling circuit and it achieved the nameplate capacity within days. The Santa Rita grinding circuit consists of one 30?x 15.4? 8.5MW SAG Mill fitted with Outotec patented Turbo Pulp Lifters (TPL TM ) in close circuit with HP400 short head pebble crusher and one 20? x 28.4? 5.5MW Ball Mill in closed circuit with Warman cyclone cluster fitted with 10 x 26? CAVEX cyclones. This paper examines the comminution design process, operating philosophies, experiences, post commissioning optimization, and future development works. Evaluation of the milling circuit performance relative to the design criteria is also discussed. Keywords: Autogenous Grate (AG), Semi-Autogenous Grate (SAG, Turbo Pulp Lifter (TPL)

Sep 1, 2012

By D. Renda

With thise Process of obtaining a solid fuel which is both ecological clean and has a high calorific value by a two-stage method of enriching preferably type lignitic coals to remove a high content of all sulphurs, incuding organic sulphur, which method comprises converting the coal to a semi-coke from followed by magnetic separation. In the Experiments given the best results the washed lignite coal in the West-Turkey, Tavsanli- Region, size -3.35 mm on dry basis. In the microscopic studies of the brigh-thin cross-section taken from the samples of sulfur min-erals pyrite in the dimension of 2~50 micrometers and melnicovite 10 micrometers particles. An thermal treatment an semi-coking themperature between 370°C to 650°C give the best Result of 480 °C. The oxidation and decomposition of sulphur minerals pyrite and melnicovite in lignite coal proceed the iron oxides Fe2O3 and pyrrhotite FeS on the surface layer. The Magnetic Seperator (Rare Earth Magnetic Separator, REMS) is capable this two form of sulphur, including organic and inorganic sulphur, in semi-coke and resulted in the magnetic-fraction in REMS. The non-magnetic-fraction remove total sulphur by 68 %, ash by 44 % and increase the upper calorific value in 15 %. The clean lignitic coal can be utilized in human or energy power stations.

Jan 1, 2014

By S. Castro

The aim of the present paper is to review the current technologies for arsenic removal in the copper metallurgical industry, and to show the progress made with some promising emerging strategies, involving the pre-treatment of copper-arsenic concentrates by selective flotation or by alkaline leaching with Na2S-NaOH solutions. The advantages of alkaline leaching are the selective dissolution of arsenic and antimony, the high extraction yield, and the practically nil loss of copper, which remains in the clean copper concentrate as a CuS residue. Alkaline ?cured? ? a procedure similar to ?acid-cured? in conventional leaching ? improves arsenic extraction and reagent consumption, making this an interesting technology for the treatment of high As-Sb concentrates. In light of recent advances on the selective flotation of enargite and tennantite, the importance of redox conditions to control the flotation/depression of enargite-tennantite from copper sulfides has been well established. However, problems of selectivity must still be solved. Additionally, the limitation imposed in some cases by the need to recover the molybdenum by-product must be considered. The application of selective flotation to obtain a high-arsenic or ?dirty? concentrate and its sub-sequent treatment by the alkaline leaching process is suggested.

Jan 1, 2014

By Silvia França, Hudson Couto

"In this paper, an alternative beneficiation process was developed to recover nickel minerals from low-grade weathered ores. The studied sample, 0.3% Ni grade, is essentially composed of nickel sulfides (pyrrhotite and pentlandite); magnetite and silicates (mostly talc and amphibolites) are the primary gangue minerals. The particle size distribution was controlled using a grinding operation to avoid producing an excess of fine particles by overgrinding. The sample underwent a grinding step for the fine degree of liberation. The slimes (-20 ?m) were removed via two desliming stages in hydrocyclones (diameters of 1.5" and 2"); the ground material, which was classified in a range of (-150+20) ?m, was processed by column flotation in three different stages: rougher flotation for magnesium silicates, rougher flotation for sulfides, and cleaner flotation for sulfides, using pilot flotation columns (2" and 6" in diameter and 20 ft high). The goal of using the preliminary magnesium silicate flotation step was to remove hydrophobic minerals (mostly talc) that would float naturally with the sulfides, which would decrease the Ni grade in the concentrate, as well as to reduce the MgO content because this compound is considered harmful in metallurgical processes when a grade greater than 9% is used. The final concentrate obtained in the cleaner flotation step reached up to 2% Ni and 7.0% MgO. The global and flotation recoveries in the process were approximately 30% and 60%, respectively, due to the losses of Ni during the desliming process. Despite this low global recovery, the adopted beneficiation process can be considered an interesting alternative to produce nickel concentrate from tailings or waste materials."

Jan 1, 2014

By S. M. H. Haddad, M. Saghaeian, R. Hejazi

"Due to limited water resources in mineral processing plants, increasing the efficiency of iron tailing dewatering system in order to obtain more usable water is very important. In this research, flocculation process, as one of the best method for water recovery, was investigated by using one organic cationic and three anionic types of flocculants with different dosages (10, 20, 30 g/ton) and iron ore tailing slurry (8 % w/w solidity) in series of graduated cylinders and obtained results were compared. it is observed that for tailing sample, the anionic flocculants have a significant effect on the settling rate of solid particles compared to the cationic types. Furthermore, effects of temperature (10, 23 and 43 ºC) and pH (2, 7 and 12) on the rate of settlement and turbidity was studied by utilizing the best flocculant. Findings showed that by increasing the amount of flocculant dosage, sedimentation rate went up and turbidity decreased. Comparison the out-put data of tests indicated that at temperature 23 °C and pH=7, the best results for both of faster settling rate and producing more clear water were reached. It should be considered that lower temperature has a more negative effect on the sedimentation rate and overflow water quality. So it will be predicted that for cold seasons, the consumption of the flocculant powder would be higher than the moderate or warm seasons. Also, coagulation/flocculation was evaluated by applying organic cationic coagulant added to anionic flocculant in three different times. Different ratios of coagulant to flocculant was examined (70%:30%, 50%:50%, 30%:70%) as well. Overall, utilizing coagulant had desirable influence on sedimentation experiments and obtaining high quality water. 1- INTRODUCTION In mineral Factories, among the important processes, solid-liquid separation process due to limited resources of water is very important and operates on iron ore tailing in settling tank as thickener in order to gain clear water for preventing water loss and applying again in plant units (Ciftci and Serhat Isik, 2017). Iron ore tailing include inharmonious compounds of iron oxides, silicates and aluminum-silicates. The percentage of solid in the tailing usually changes between 5 and 10 percent (Dash et al., 2011). One of the major problem about dewatering of tailing in thickener is solid-liquid separation of slurries with high stability because of fine particles. It means that rate of sedimentation is very slow and recovered water quality is inappropriate. This issue can be solved to a large extent by factors such as pH control, ionic strength and flocculation/coagulation process which results in aggregation of particles so that settlement by gravity is happened (Dash et al., 2011)."

Jan 1, 2018

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

"Impact breakage is the most elementary size reduction mechanism in comminution devices. In comminution research, characterization of breakage is primarily concerned with the energy expended to cause it. The main challenges of this approach have been quantifying the actual energy requirements of impact breakage and developing mechanistic models to accurately describe the process. For engineering load analysis, fracture is characterized in terms of strengths such as yield stress and ultimate tensile stress, which govern the maximum loads specific materials can be subjected to before irreversible inelastic deformation and subsequently complete fracture occur. With this approach, the main challenge has been to elucidate the underlying mechanism, in the means that the load causes weaknesses in the material microstructure to act as sites for stress concentrations, leading to crack initiation, propagation and eventual failure. Studies dedicated to investigating particle fracture under impact loading have led to the development of several devices for conducting breakage characterization tests. These devices include the drop weight tester, split Hopkinson pressure bar, rotary breakage tester and the short impact load cell. All these have different particle loading and breakage methods, and thus can yield distinct information which can be compiled together to study a material and the nature of breakage. This work analyses the single impact breakage data obtained across the listed devices. A homogenous ore commonly known as blue stone is used for all breakage tests. The work is conducted using standard procedures specific to each device as closely as possible. Drop weight and rotary breakage tests were used to obtain hardness parameters based on an existing t10 breakage model, while split Hopkinson pressure bar and short impact load cell measurements provided properties based on conventional engineering load analysis. The results were used to highlight the significance of the loading geometry, impact mechanism and impulse on the extent of breakage."

Jan 1, 2018

By John A. Rumball, Martin R. Houchin

Data from autogenous and semi-autogenous tests in a 1.67m diameter mill have been analysed in terms of a new model for rock abrasion. This model defines specific rate as the rate of production of fines (material less than 13mm) per unit mass of rock, where the term “rock” is applied to all ore pieces of ore coarser than the grate size. In most cases, an allowance for preliminary rounding and cleavage was made. A method has been developed to analyse steady-state data, using piecewise fitting to the cumulative size distributions. The model was compared to the well-established models, which are based on the population balance method. The specific rate increased for all rock sizes as the volume of 100mm balls in the mill was increased. Tests using a coarser feed and smaller balls provided further evidence that once rocks are rounded, abrasion is due to localised surface damage, caused by impacts, rather than rotational shear. An allowance for rapid rounding of jagged rocks, followed by less rapid abrasion of rounded pebbles was also tested, but the analysis was not conclusive. The average specific rate is a useful measure of the capacity of a primary mill. The effect of power intensity (kW/t of ore) was also investigated as a means of extrapolating the pilot-plant data to large production mills.

Sep 29, 2003

By Xiao Ni

Alkyl hydroxamic acid was shown to be a suitable collector for the flotation of pyrochlore from carbonatite niobium oxide ores. In the direct batch rougher flotation from the ore without desliming, over 85% of the pyrochlore can be recovered at about 30% mass pull. Cleaner flotation tests showed that the rougher concentrate can be upgraded. The adsorption mechanisms of octyl hydroxamic acid (OHA) were studied by using high purity pyrochlore and calcite mineral samples and employing analytical measurements such as Fourier transform infrared spectroscopy (FTIR), X-ray photoelectron spectroscopy (XPS) and UV/VIS spectroscopy. FTIR showed a shift of the vibrational frequencies of the carbonyl group in hydroxamic acid after adsorption on pyrochlore and calcite, and the shift was more pronounced on pyrochlore. XPS results indicated a lowering of the binding energies (chemical shift) of Nb3d5/2 and Ca2p3/2 electrons on pyrochlore, and the lowering of the binding energies of Ca2p3/2 electrons on calcite, again with the shift being more pronounced on pyrochlore. The results indicated hydroxamic acid formed stronger complexes with pyrochlore than with calcite. Adsorption isotherms indicated that OHA adsorbed more on pyrochlore than on calcite at low concentration. At high concentrations, the adsorption densities on both minerals were similar. All results explained the observed selectivity of the reagents in the flotation of the carbonatite niobium oxide ores.

Sep 1, 2012

By W. Krieglstein

Due to the natural variability of ores, it is necessary to carefully test flotation behavior of flotation cells before designing flotation circuits. For performing quick and versatile tests in the laboratory, Siemens developed a laboratory-size Hybrid Flotation Cell, which has been down-scaled from the 300 m³/h production cell to a 0.6 m³/h flotation cell. This small cell is being used as a test bed for technical and chemical concepts, the improvement of decisive components and the characterization of customer ores. In this study the evaluation results for Andean porphyry copper ores are shown. Total recovery rates of the order of 45 to 50 % have routinely been achieved in a single flotation cycle over a wide range of feed grades. In addition, the Hybrid Flotation Cell has proven to be particularly well suited for the flotation of fine particles below 20 µm down to below 1 µm, at high recovery rates and concentrate grades.

Sep 1, 2012

By A. Ghadyani

This investigation was undertaken to assess the possibility of ash and sulfur removal by microbial action of high sulfur-content coal. Thiobacillus ferrooxidans and Thiobacillus thiooxidans have been used for biodesulfurization of coal. Feed samples were provided from Tabas coal mine located in Yazd province, Iran. The sample contained 28.80% ash and 2.30% sulfur. Flotation tests were implemented to optimize the effective parameters in these cells which were constant during bio-flotation tests. According to the results, obtained from various tests using Taguchi orthogonal array design (L27), the values of optimized parameters were pH at 7, 1000 g/t gas oil as a collector, 10% solid content, 15 min conditioning time, 100 g/t MIBC as a frother and 800 rev/min rotor speed. Bio-flotation tests are carried out by means of laboratory double discharge cell and mesophilic bacteria as depressant reagent using above mentioned optimized parameters. The bio-flotation experiments were conducted based on a Taguchi orthogonal array design (L9) with four variables (bacteria type, bacteria count, rotor speed of mixer and bio-conditioning time). The collected data was analyzed by using DX7.0, Minitab14.0 and WinRobust1.02 softwares for evaluation the effect of each parameter on the optimization criteria. Considering the highest combustible recovery and the lowest ash and sulfur content, the value of optimized bacteria type, bacteria count, rotor speed of mixer and bio-conditioning time were Thiobacillus ferrooxidans, 5×107 cells/g(coal), 100 rev/min and 20 min respectively. Under this condition, ash and sulfur content and combustible recovery were predicted 12.39%, 2.08% and 82.03% respectively.

Sep 1, 2012

By Yi Ran Zhang, Nawoong Yoon, Brent Hilscher, Maria Holuszko

"Ore sorting technology has been rapidly advancing in recent years. Improving expected mill feed grade has a significant benefit to project returns, and can no longer be ignored. When there is an increase in the mill feed grade while maintaining the same throughput, the mine will be able to process much more metal and improve the floatation recovery. There are many more factors that contribute toward the economics of the project by implementing sensor-based sorting technologies covered in this paper. This paper reviews the economic impact on an underground copper-silver deposit with references to the concept of ""profit recovery,"" which is the measurement of recovery based on rock profitability rather than just metal content. The economic impact of sensor-based sorting is discussed in detail throughout the paper. The preliminary economic evaluation of the project compares the mine's base case design versus a design incorporating sensor-based sorting. The paper's primary focus is on economic impacts rather than the explanation of the sensor based sorting results. The sorting data is presented on a scoping level and using a sensor-based sorting study where the best possible results are used for obtaining the grade recovery curves. Also, the breakdown and analysis of the project economics are based on the operating cost available from the particular operation. Any assumptions taken into consideration are also outlined and discussed. INTRODUCTION The mining industry has been adopting Sensor-based ore sorting technology fairly rapidly in the recent years. The ability to increase mill feed grade by rejecting waste grade rocks early in the processing stage is proving to be revolutionary for some operations. These potential outcomes should not be overlooked by both producing operations and prospective projects, thus leading to a wide range of studies on ore sorting. The answer to this is the creation of a new term and concept - Profit Recovery™."

Jan 1, 2018

By A. Norori-McCormac

"Air recovery, the fraction of air entering a flotation cell that overflows the cell lip, is linked to flotation performance, and it has been shown through industrial testwork that operating flotation cells at air rates that yield the Peak Air Recovery (PAR) results in high mineral recovery. For operating parameters over which flotation plants have little control, such as particle size, investigating the effect on the position of PAR, and the subsequent effect on flotation performance, proves more challenging. To this end, we have developed a bench scale flotation system that runs continuously and exhibits froth behavior similar to that found industrially, allowing the study of such parameters as particle size in addition to air rate. The bench scale system comprises a 4 l baffled cylindrical flotation cell, stirred by a Rushton impeller. The concentrate is recycled back into the pulp, allowing for continuous operation. Glass beads are used as solid particles, allowing a wide range of particle sizes to be tested, with TTAB as collector and MIBC as frother. The system allows operation at air rates typical of industrial flotation systems (1 -2 cms-1 superficial gas velocity), in addition to yielding a froth that coalesces and bursts and exhibits a peak in air recovery as cell air rate in increased. Solids and water recovery is also measured. The results show that air recovery is sensitive to changes in particle size. The intermediate size distribution showed a peak in air recovery (PAR) that corresponded with the maximum solids flowrate, as found industrially. It is shown that finer particles do result in more stable froths and higher recoveries, particularly at low air rates. The effect of particle size on the position of PAR is not yet unequivocally determined from these results."

Jan 1, 2014

By E. Burdukova

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

Jan 1, 2014

"A sulfide ore containing 2.19 g/t Au and 13.13 g/t Ag was studied for the deportment and extractive metallurgy of pay metals. The ore is composed mainly of non-sulfide minerals with approximately 10% pyrite and minor amounts of marcasite, chalcopyrite, galena, sphalerite, and arsenopyrite. The deportment study using a comprehensive approach showed that gold and silver minerals in the ore are mainly native gold (Au), electrum (Ag, Au), petzite (Ag3AuTe2), hessite (Ag2Te) and volynskite (AgBiTe2) with minor amounts of acanthite (Ag2S) and argyrodite (Ag8GeS6). At current grinding fineness (P70@150µm), the majority of liberated native gold and electrum particles are larger than 20µm with 85% >40µm, some of them are even larger than 150µm, indicating that extended leach time may be required to dissolve these coarse gold grains to completion and gravity concentration needs to be considered. Petzite is relatively fine with all liberated particles being less than 80µm in size. Gold distribution and leach testwork results showed that liberated gold accounts for 50% of the head assay, exposed gold associated with sulfide and non-sulfide minerals accounts for 30% and 11% of the head assay, respectively. Gold locked in sulfide and non-sulfide accounts for 4.8% and 4.1%, respectively. Thus, it is recommended that cyanide leaching be considered for extracting gold in this ore. This paper presents the detailed results of the gold and silver deportment study and extractive metallurgy testwork, and discusses the impact on process selection and flowsheet development.INTRODUCTION Gold and silver have been found in various geological environments and ore types. In terms of the gold deportment and extractability, gold ores can be classified into two categories and 12 types. Their extractive metallurgy is largely driven by mineralogical factors such as particle size, association with other minerals, coatings and rimmings, presence of cyanicides, oxygen consumers and preg-robbers, presence of refractory gold minerals and locking of submicroscopic gold in sulfide and sulfarsenide mineral structures (Zhou et al., 2004; Marsden & House, 2006; Zhou, 2013a, 2013b). Compared to gold, silver mineralogy is more complex and metal extraction is more challenging due to its presence in over 200 Ag-bearing minerals (a silver-rich ore body would typically contain 5 to 10 different silver minerals) (Gasparrini, 1984), presence of refractory silver minerals (such as pyrargyrite, prustite and stephanite) and submicroscopic silver in sulphide minerals (such as galena and copper minerals) (Chryssoulis et al., 1985; Di Prisco et al., 2002). The presence of these issues often causes problems in Ag-bearing ore processing. Therefore, knowing the deportment of gold and silver in the ores and metallurgical products is important for flowsheet development and process optimization."

Jan 1, 2018

By Brian Hart, Rong Fan, Roger Smart, Andrea Gerson, Ning Xu

A complete strategy has been developed for identification of the reasons for losses in recovery and/or grade, due to changes or complications in mineralogy and in mineral surface conditioning, and their subsequent correction. The sequence of methods starts with comparison of mineral phases between QEMSCAN/MLA (compositional) and Rietveld XRD (crystalline diffraction) with bulk assay reconciliation. This methodology has recently led to identification of unrecognised amorphous content (12-40 wt.%), previously attributed to crystalline phases, in the fine fraction (<35 µm) of some problematic flotation feeds. Attached amorphous silica, talc- and chlorite-like fines have direct influence on the hydrophobic/ hydrophilic ratio on value mineral surfaces adversely affecting flotation. SEM/EDS is used in the backscattered imaging mode to determine possible problems with liberation (e.g. remnant local locking), not detected in standard liberation analysis, resulting in gangue flotation. EDTA extraction is used to assess bulk feed pre-oxidation and sequential oxidation product loading through the circuit. Changes in circuit solutions are assayed with speciation modelling of precipitates and potentially adsorbing species. ToF-SIMS with principal component analysis is then used to correlate surface species with specific minerals and can be used to assess the effectiveness of specific reagents particularly collectors. This approach identifies unwanted activating species and interference with value surfaces by adsorbed hydrophilic ions, precipitates and fine particles. When the interference or ineffective reagents are specifically identified, remedial action can be targeted. This method has identified residual talc-like layers (<10 nm thick) inducing gangue flotation of pyroxene and chromite in PGM flotation. Recent case studies of application of the full methodology to losses of molybdenite at KUCC and to sphalerite activation in the Cu/Pb circuit at Brunswick Mines with remedial action are discussed.

Jan 1, 2014

By Xinyang Xu, Chen Zhang, Lixia Li, Zhitao Yuan

"Density functional theory was employed to clarify the mechanisms of the activation of a quartz surface and its collection by the flotation method at the molecular level. The calculation results indicated that hydrogen bonds on the quartz surface (1 0 1) existed in pairs while in contact with water. The simulation was carried out to investigate the different effects of Ca2+ and Ca(OH)+ on a quartz surface with hydrogen bonds, and the results suggested that Ca(OH)+ most likely adsorbed onto and activated the quartz surface. Mulliken population analysis of the external oxygen atom of the quartz surface, calcium atom of Ca(OH)+, and the oxygen atom of oleate showed that there were electron transfer and the formation of a chemical bond between Ca1-O1 and Ca1-O3, which were indicative of strong interactions between the three atoms. In addition, the –OH of the Ca(OH)+ contributed to the activation, which suggested that the Ca(OH)+ group acted as a bridge between the surfaces of oleate and quartz.INTRODUCTIONFlotation is an effective separation method in mineral processing that takes advantage of the physicochemical differences between minerals and gangue, and the properties of the mineral surfaces play a key role in this process (Liu et al. 2016, Qin et al. 2017). In the reverse flotation process of iron ore, gangues are captured by cationic or anionic collectors and floated with air-bubbles as tailings, and the pre-depressed iron oxide minerals are left at the bottom as concentrates (Lima et al. 2013, Yianatos et al. 2015, Zhu et al. 2015). Quartz (SiO2) is one of the major gangues of hematite ores and cannot be easily removed. It is a naturally hydrophilic mineral, in which silicon is the cation in its crystal structure and has very limited solubility. Therefore, quartz must be activated, usually by metal ions, for its effective rejection as a float-product by reverse flotation (AIME. et al. 1976)."

Jan 1, 2018

By Lucy van de Ruit, Megan Becker, Aubrey Mainza, Jenny Wiese

"UG2 chromitite ore is a complex, highly disseminated platinum ore from the Bushveld Complex in South Africa, with typical platinum group mineral (PGM) grain sizes below 15µm. The introduction of stirred mills to the platinum industry has enabled grinding to finer product sizes, which has improved liberation and recovery of the finely grained PGMs. However, reduction in particle size leads to an increase in chromite entrainment in the flotation concentrate, which has a negative impact on the subsequent smelting process. Factors such as particle size, froth structure, solids concentration and water recovery have all been identified as impacting entrainment, but whether particle shape also has a significant effect is unclear. The purpose of this study was to determine whether the recovery of chromite particles by entrainment during the flotation of UG2 ore is affected by particle morphology. A laboratory vertical stirred mill with inert ceramic media was used for grinding UG2 ore which was then floated in a batch flotation cell. Flotation feed and concentrate samples were analyzed using QEMSCAN to quantify the angularity and aspect ratio distributions of the particles. The stirred mill product was found to have relatively broad aspect ratio and angularity distributions, with the major mineral components (chromite and silicates) having very similar shape characteristics. Analysis of flotation feed and concentrate samples suggested that particle shape does not have a significant effect on chromite entrainment."

Jan 1, 2014