Sulfide Flotation Research Articles

Flotation separation of copper–molybdenum sulfides using chitosan as a selective depressant

Chitosan, a natural biodegradable polymer, was studied as a selective depressant in copper–molybdenum sulfide separation by flotation. Although chitosan depressed both molybdenite and chalcopyrite in single mineral flotation, selective flotation was achieved during the flotation of molybdenum-bearing copper sulfide concentrate at pH 6; the recovery of molybdenum in the froth product was over 70% while the recovery of copper was less than 24%. Adsorption measurements showed that chitosan had a higher adsorption density on chalcopyrite than on molybdenite. FTIR and ToF-SIMS measurements were carried out to confirm the different degrees of interaction of chitosan with molybdenite and chalcopyrite and the results showed that chitosan mostly adsorbed on chalcopyrite via amide groups when both molybdenite and chalcopyrite were present in the suspension.

Contribution of chloride ions to the sulfidization flotation of cerussite

The effect of the addition of chloride ions on the sulfidization flotation of cerussite was investigated by micro-flotation tests, surface adsorption experiments, zeta potential measurements, and X-ray photoelectron spectroscopy (XPS) analysis. The micro-flotation test results indicated that the addition of chloride ions prior to sulfidization significantly increased the flotation recovery of cerussite. The maximum and minimum percentages of increase in flotation recovery were 16.63% and 10.51%, respectively. Surface adsorption experiments of sulfide demonstrated that there was a higher adsorption concentration of S species on the mineral surface, and its adsorption efficiency was enhanced. The results of zeta potential measurements revealed that the addition of Na2S and chloride ions to the solution caused a more negative zeta potential than the addition of Na2S alone. The XPS analysis results indicated that more lead sulfide formed on the mineral surface after the addition of chloride ions into the pulp solution. These results consistently showed that the addition of chloride ions played a significant role in the sulfidization of cerussite, resulting in a better flotation performance.

Optimized activity ratio for different types of reagent attachment at sulfide minerals

The authors evaluate the ratio of activities exhibited by physical and chemical adsorptions of an agent on sulfide minerals. The following assumptions are made: 1) the main kinetic restriction for flotation contact (particle–bubble interaction) is the film between mineral particle and air bubble; 2) water is removed from the film by both physical and chemical adsorptions; 3) the water volume removed by physical adsorption is conditioned by the difference in the volumes of water in the films bounded by meniscuses with advancing angle ΘA and receding angle ΘR and neck radius R. The water volume removed under the action of surface forces (hydrophobic, dispersive) was disregarded as the agent attachment was low dense and mosaic. The article reports experimental data on limit values of static advancing and receding angles on the surface of some sulfide metals. It is shown that selectivity of flotation of sulfides depends on the ratio of water volumes removed from the film by desorbable physical adsorption and non-desorbable chemical adsorption types of agents. The proposed method of selectivity estimation is suitable for minerals with similar (close) properties of the surface when stabilizing ion–electrostatic interaction between minerals and air bubbles is absent or insignificant. The calculation results may be of use to optimizing collecting ability of physically adsorbed low-polar or desorbable species of ionized agent with intent to improve selectivity of separation of sulfide minerals.

Contribution of components from volume defect in natural pyrite and quartz to solution chemistry of flotation pulp

The volume defects in pure pyrite and quartz from a classical Cu–Pb–Zn–Fe sulfide deposit were investigated. The results indicate that a large number of volume defects exist in natural pyrite and quartz. The volume defects assume a variety of shapes, including long strips, oval shapes and irregular shapes, with sizes ranging from a few microns to dozens of microns. These volume defects are rich in metallogenic elements as a result of the capture of metallogenic and mineralizing fluid during the defect-forming process. The volume defects are fractured during the grinding process, and their chemical components are released into the solution, as confirmed by the abundant presence of various metal and non-metal components in the cleaning water and EDS results. Under the experimental conditions of 10 g pyrite or quartz with grinding fineness of d90=37 μm, which was cleaned in 40 mL of pure deionised water under an inert atmosphere, the total average concentrations of Cu, Pb, Zn, Fe, Ca, Mg and Cl− in the aqueous solution are 32.09×10−7, 16.51×10−7, 19.45×10−7, 516.52×10−7, 129.50×10−7, 35.30×10−7 and 433.80×10−7 mol/L, respectively, for pyrite and 19.20×10−7, 8.88×10−7, 8.31×10−7, 82.71×10−7, 16.21×10−7, 4.28×10−7 and 731.26×10−7 mol/L, respectively, for quartz. These values are significantly greater than those from the experimental non-oxidative dissolution of the pyrite or quartz, respectively. Therefore, the metallogenic fluid in volume defects of mineral crystal is concluded to represent the dominant contribution to the solution chemistry of sulfide flotation pulp. The present investigation will help to deeply understand the flotation theory of sulfide minerals.

카자흐스탄 구리제련소 슬래그 내 구리회수를 위한 선별 특성

Abstract Almost all copper slags contain a considerable amount of Cu (0.5 - 3.7%) close to or even higher than copper ores. A numberof methods for metal recovery from copper slag were reported These methods can be classified into three categories, flotation,leaching and roasting. Sulfide flotation method for the recovery of copper from Kazahstan copper furnace slag is discussed inthis investigation. 50% of copper from the slag was recovered by sulfide flotation at pH 4. meanwhile 67% of copper from theslag was recovered at pH 11. Higher copper recovery result at pH 11 rather than that at pH 4 was caused by the fact that coppersulfides were floated in particle size fraction over 100µm in concentrates at pH 11. When the slag were ground below 74µmby ball milling, the recovery of copper by floation in slag improved to 78 - 83% because of copper liberation effect.Key words : slag, copper slag, flotation, magnetic separation, recovery of copper 1. 서 론 구리제련 슬래그는 광산에서 선광과정을 거쳐 발생한구리정광을 고온에서 녹여 조동(anode)을 생산하는 과정 중, 융제인 규사가 철과 결합하여 발생하는 부산물로 구리가 주성분인 매트(matte)와 분리 배출된다.

Flotation of copper sulphide from copper smelter slag using multiple collectors and their mixtures

Present work focuses on the differences in the performances obtained in the froth flotation of copper smelter slag with multiple collector viz. sodium iso-propyl xanthate (SIPX), sodium di-ethyl dithiophosphate (DTP) and alkyl hydroxamate at various dosages. Flotation tests were carried out using single collectors as well as various mixtures of the two collectors at different but constant total molar concentrations. Flotation performances were increased effectively by the combination of collectors. The findings show that a higher copper recovery (84.82%) was obtained when using a 40:160g/t mixture of sodium iso-propyl xanthate (SIPX) with di-ethyl dithiophosphate compared to 78.11% with the best single collector. Similar recovery improvement (83.07%) was also observed by using a 160:40g/t mixture of sodium iso-propyl xanthate (SIPX) with alkyl hydroxamate. The results indicate that in both cases DTP and alkyl hydroxamate played important role as co-collector with SIPX for the recovery of coarse interlocked copper bearing particles and has an important effect on the behaviour of the froth phase. However, comparatively higher mass recovery was obtained when hydroxamate was used as a co-collector, accompanied by poorer copper grades. In this context the co-collector DTP is advantageous for the improvement of copper grade and recovery from copper smelter slag.

Composition of multicomponent heracleum extracts and its effect on flotation of gold-bearing sulfides

Using the proprietary method, the authors have obtained flotation agent BO—extract of atomized leaves and stalks of Heracleum sosnowskyi (hogweed), and analyzed its composition. Qualitative tests, thin-layer chromatography (TLC), ultraviolet spectrophotometry and flotation tests show that aqueous and organic extracts contain substances that differ in chemical and process properties. The article gives test results for aqueous, alcoholic and aqueous–alcoholic extracts BO used as modifiers in chalcopyrite and pyrite flotation. Selective properties of aqueous and aqueous–alcoholic extracts BO are characterized. The highest depression ability relative to pyrite is a characteristic of aqueous BO extract obtained at pH = 3, with heating and treatment in ultrasonic bath. Organic extracts contain mostly furocoumarins, aspic oils and resins, and have no material effect on flotation activity of pyrite.

SELECTION OF SULFHYDRIL COLLECTORS FOR FLOTATION OF SULFIDES OF NONFERROUS METALS FROM REFRACTORY ORES OF NONFERROUS METALS

Methods of selection of selective collectors are analyzed. The procedure of calculation of characteristics of absolute rigidity, absolute electronegativity, and reaction ability of compounds is presented. The correspondence between the results of theoretical calculations of the reaction ability of sulfhydril collectors and experimental results of adsorption and flotation of monomineral fractions of chalcopyrite, galenite, and pyrite by sulfhydril collectors is revealed. The prediction of the collecting activity based on calculations is complicated by the features of the chemical composition of natural sulfides and defectiveness of the surface associated with the genesis of ore minerals in the deposit is revealed. Dithiophosphates and thionocarbamates are isolated as main components of selective compositions of collectors for flotation of pyrite ores of nonferrous metals based on theoretical and experimental investigations. It is shown by the complex of adsorption investigations in conditions of nonfrothing flotation and analysis of the IR frustrated total internal reflection (FTIR) spectra of concentrates of nonfrothing flotation that the nonoptimal ratio of components weakly active relative to pyrite leads to a considerable increase in pyrite floatability. Sorption of components of the sulfhydril collector and pyrite floatability increase with the fraction of a nonionogenic collector in the composition lower than 40–60%. The action mechanism, which determines the action selectivity of sulfhydril collectors during the flotation of sulfides, is determined.

Влияние поверхностно-активных свойств реагентов, содержащих диалкилдитиофосфаты натрия, на флотацию сульфидов

Влияние поверхностно-активных свойств реагентов, содержащих диалкилдитиофосфаты натрия, на флотацию сульфидов

Selection of sulfhydryl collectors for flotation of sulfides of nonferrous metals from refractory ores of nonferrous metals

Methods of selection of selective collectors are analyzed. The procedure for calculating characteristics of absolute rigidity, absolute electronegativity, and reaction ability of compounds is presented. The correspondence between the results of theoretical calculations of the reaction ability of sulfhydryl collectors and experimental results of adsorption and flotation of monomineral fractions of chalcopyrite, galenite, and pyrite by sulfhydryl collectors is revealed. The prediction of the collecting activity based on calculations is complicated by the features of the chemical composition of natural sulfides, and the defectiveness of the surface associated with the genesis of ore minerals in the deposit is revealed. Dithiophosphates and thionocarbamates are isolated as main components of the selective compositions of collectors for the flotation of pyrite ores of nonferrous metals based on theoretical and experimental investigations. It is shown by a complex of adsorption investigations in conditions of nonfrothing flotation and analysis of the IR frustrated total internal reflection (ATR) spectra of concentrates of nonfrothing flotation that the nonoptimal ratio of components weakly active relative to pyrite leads to a considerable increase in pyrite floatability. The sorption of components of the sulfhydryl collector and pyrite floatability increase when the fraction of a nonionogenic collector in the composition is lower than 40–60%. The action mechanism, which determines the action selectivity of sulfhydryl collectors during the flotation of sulfides, is determined.

INVESTIGATIONS OF GOLD BEARING ORES BENEFICATION

There were quoted results of sulphide flotation with regard to ores of Natalka deposit. There was showed a possibility of production from ore contended 2 g/t of Au concentrate with gold content from 40 to 95 g/t at release of it over the range from 30 to 5 %. Precious metal recovery in such a case is 90–35 %. Sorption leaching of received concentrate by virtue CIL method allows recovers from it from 87,1 to 96,4 % of gold that is equivalent to total recovery in the range of 75,2–83,2 %.

RECOVERY OF COBALT AND NICKEL BY ATMOSPHERIC LEACHING OF FLOTATION SULFIDE CONCENTRATE FROM LUBIN CONCENTRATOR

An experimental study on leaching of cobalt and nickel from a Lubin (Poland) sulfide flotation copper concentrate with oxygenated sulphuric acid solution in the presence of iron(III) sulfate is presented. The effect of temperature, iron(III) concentration, initial sulphuric acid concentration, oxygen flow rate and chloride ions addition on cobalt and nickel leaching were examined. The obtained results show, that enhancing investigated leaching parameters increases both the rate and leaching recovery of nickel and cobalt. The effect of iron(III) ions concentration is particularly significant. Additionally, a significant correlation between copper and nickel/cobalt leaching recovery was observed. Due to the fine dissemination of nickel and cobalt, they were leached remarkably faster when copper leaching recovery exceeded 90%.

An investigation into electrochemical interactions between platinum group minerals and xanthate: Voltammetric study

In the flotation of platinum group minerals (PGMs) containing significant amounts of tellurides and arsenides it is generally assumed that these minerals will respond in the same way as sulphides to collectors such as xanthates which are typically used in sulphide flotation. This paper presents the results of a fundamental study which has been conducted to provide a greater insight into the comparative interactions between synthetic moncheite (PtTe2) and cooperite (PtS) with sodium ethyl xanthate (SEX). Cyclic voltammetry has been used to investigate these interactions in the absence and presence of sodium ethyl xanthate (SEX) in aerated and de-aerated solutions. Reduction potentials of the SEX/dixanthogen couple have been measured and compared to published data. Reduction potentials of possible reactions of the minerals have been predicted from thermodynamic calculations and used to attempt to interpret the potentials measured in the cyclic voltammetry investigations. Results have shown differences in the oxidation behaviour of the minerals as well as in the SEX interaction with the minerals. The effect of pH was investigated in the case of the Pt minerals and fractional surface coverages were calculated with a view to ultimately correlate these results with flotation behaviour. Proposals are made with respect to the various reactions occurring under the conditions studied.

Flotation of Sulfide Components of Copper Ore in the Presence of n-Dodecane

Influence of n-dodecane aqueous emulsion on flotation of sulfide minerals present in the copper ore from the Legnica-Glogow Copper Basin (LGOM), carried out on the laboratory scale, was investigated. Five flotation experiments performed in the presence of different doses of n-dodecane were conducted and compared. Next, mineral analysis of selected flotation products were conducted and discussed. The best flotation selectivity was obtained for bornite, tennantite and chalcocite using 600 g/t of n-dodecane. In the case of non-copper sulfides, the maximum selectivity was observed for sphalerite and galena. Measurements of the rest potential of copper sulfide conducted to explain the differences between recoveries of each sulfide mineral were performed. It was observed, that with increasing sulfide rest potential, the recovery was decreasing. The study provided information hitherto poorly available in literature on flotation of sulfides from the LGOM area with n-dodecane.

Study on the Beneficiation of an Oxidised Copper Ores in Yunnan Province

For a refractory oxidised copper ores in Yunnan province, studing the beneficiation of a oxidised copper ores sulfide flotation behavior. The results showed that: When ethylenediamine phosphate was added with the dosage of 120 g/t, the final copper concentrates were gained with the grade of 21.06% and recovery of 83.20%. Compared to the direct sulfide flotation, the concentrate grade was raised by 3.59% and the recovery was increased by 7.33%.

Optimization of Cu–Zn Massive Sulphide Flotation by Selective Reagents

Selective floatation of base metal sulphide minerals can be achieved by using selective reagents. Sequential floatation of chalcopyrite–sphalerite from Taknar (Iran) massive sulphide ore with 3.5 % Zn and 1.26 % Cu was studied. D-optimal design of response surface methodology was used. Four mixed collector types (Aer238 + SIPX, Aero3477 + SIPX, TC1000 + SIPX and X231 + SIPX), two depressant systems (CuCN–ZnSO4 and dextrin–ZnSO4), pH and ZnSO4 dosage were considered as operational factors in the first stage of flotation. Different conditions of pH, CuSO4 dosage and SIPX dosage were studied for sphalerite flotation from first stage tailings. Aero238 + SIPX induced better selectivity for chalcopyrite against pyrite and sphalerite. Dextrin–ZnSO4 was as effective as CuCN–ZnSO4 in sphalerite–pyrite depression. Under optimum conditions, Cu recovery, Zn recovery and pyrite content in Cu concentrate were 88.99, 33.49 and 1.34 % by using Aero238 + SIPX as mixed collector, CuCN–ZnSO4 as depressant system, at ZnSO4 dosage of 200 g/t and pH 10.54. When CuCN was used at the first stage, CuSO4 consumption increased and Zn recovery decreased during the second stage. Maximum Zn recovery was 72.19 % by using 343.66 g/t of CuSO4, 22.22 g/t of SIPX and pH 9.99 at the second stage.

Studies on gold ore concentration characteristics

Results obtained during the sulfide flotation of ores from the Natalka deposit are presented. The possibility of using ore containing 2 g/t Au to obtain a concentrate with the content of gold ranging from 40 to 95 g/t and yield of 30-5% is shown. Precious metal recovery is 90-35%. Carbon-in-leaching (CIL) of the obtained concentrate allows one to extract from 87.1 to 96.4% Au, which is equivalent to 75.2–83.2% provided by through recovery.

Optimization and mechanism of gold-bearing sulfide flotation

In order to overcome the difficulty of extracting gold from gold-bearing sulfide ore by cyanide process, flotation was adopted based on mineralogical analysis. Mineralogy shows that gold particles are of superfine structure and mainly enclosed by sulfide ores. Primary gold-bearing sulfide ore is fine-grained pyrite and arsenopyrite. The paper describes the effects of ratios and dosage of activators and collectors on the recovery and grade of gold concentrate. A proper flotation flowsheet was then proposed based on experimental condition and closed-circuit test. The gold concentrate with the gold grade of 25.14 g·ton−1 and the recovery of 86.94 % is obtained after one rougher, three cleaners, and four scavengers from fine grinding flotation process. Furthermore, the mechanisms of combined activators and combined collectors were studied by thermodynamic calculation, and structure–activity relationship of flotation reagent was also explained.

Designing mine tailings for better environmental, social and economic outcomes: a review of alternative approaches

Abstract With the unprecedented rate of mining and mineral processing over the past decade, there has been an increasing level of concern worldwide about the fate of tailings and the irreversible consequences of contaminant release through dust, tailings dam seepage, dam wall failure, or direct disposal of tailings into waterways. It is likely that in the future the mining of higher tonnages of lower grade ore and the possibility of more extreme weather events associated with global climatic change, may increase the risks associated with mine tailings management. In this paper we review alternative tailings management methods that aim to improve environmental, social and economic outcomes. Three approaches in particular are highlighted: paste and thickened tailings; tailings reuse, recycling and reprocessing; and proactive management (e.g. the integration of sulphide flotation with cemented paste backfill). The aim of the present study is to scope the opportunities that could be gained by developing an integrated research framework that traces tailings properties back to the orebody. In the future models could be constructed to optimise environmental, social and economic outcomes of tailings management across the value chain through integrated orebody characterisation, mine planning, processing, disposal, re-processing, recycling and reuse.

The effect of non-sulphide gangue minerals on the flotation of sulphide minerals from Carlin-type gold ores

Flotation pre-concentration of sulphide and gold values from certain Carlin-type deposits characterised as double-refractory gold ores is quite challenging. Numerous studies conducted on these ores in many laboratories globally (including the present study) under a variety of chemical and physical conditions have merely confirmed low recovery (and poor concentrate grades) for sulphide minerals and gold, and poor separation between sulphide minerals and carbonaceous matter, even when the valuable minerals are adequately liberated. None of the traditional reasons based on liberation or the choice of chemical and physical conditions and separation strategies could provide satisfactory explanation for the observed poor separation.In this study, the focus was on the role of non-sulphide gangue (NSG) minerals. It was hypothesised that specific NSG minerals have a detrimental effect on flotation recovery of gold bearing minerals and their separation selectivity. In order to test this hypothesis and delineate the effect of the various gangue minerals, a new approach was taken. This involved first isolating the various mineral components of a double-refractory gold ore from one of the Carlin-type deposits using a gravity separation technique. Then flotation experiments were performed using a mixture design on various mixtures of these isolated components under controlled conditions. The results of these mixture experiments supported the hypothesis and demonstrated, for the first time for these types of ores, that even small amounts of NSG minerals, especially carbonaceous matter and clays, had a large adverse effect on the flotation of sulphides and selectivity of separation. While it is tempting to attribute the observed effects solely to slime coating, there is no basis to do so at this stage; it is more reasonable to propose that multiple contributions exist. The results of this study provide the much-needed context and direction for further fundamental studies and for developing processing strategies.