No full-text available
To read the full-text of this research,
you can request a copy directly from the authors.
Rejection of antimony and bismuth in sulphide flotation – a literature review
Abstract
Rejection of deleterious impurity minerals containing antimony and bismuth from sulphide deposits by flotation has been reviewed. The key antimony minerals of commercial interest to remove are stibnite, tetrahedrite, and jamesonite, while the main bismuth mineral is bismuthinite. The most promising methods for separating antimony and bismuth minerals from sulphide ores are based on pH and/or pulp potential adjustments or the use of starvation levels of collector. Stibnite has been floated with xanthate, dithiophosphates and thionocarbamates in acidic conditions, but floatability decreased with increasing pH. Antimony minerals are activated with lead and copper salts and depressed with an oxidant (e.g. H2O2 or Na2Cr2O7). Bismuth minerals can float strongly between pH 3 and 7 with a xanthate collector but floatability decreased sharply at pH values above pH 7. Molybdenite is separated from bismuth minerals by depressing them with reducing agents such as sodium sulphide, while cyanide depresses chalcopyrite.
... Sb is usually considered a penalty element in refractory gold ore and copper concentrates. The presence of Sb decreases the gold extraction rate (Celep, Alp, and Deveci 2011a) and increases the cost of copper smelting and refining operations (Lane et al. 2016;Smith, Bruckard, and Sparrow 2022). The separation or recovery of Sb from the tetrahedrite and refractory gold ore is also beneficial to downstream copper smelting and gold recovery (Bhappu 1990;Filippou, St-Germain, and Grammatikopoulos 2007), respectively. ...
The global reserve of antimony is gradually declining while its demand continuously grows, rendering antimony an increasingly scarce metal. Antimony has been highlighted as a critical raw material in the EU owing to its economic significance and high supply risk. Due to the dispersed uses and low antimony concentration of the main antimony end-use, i.e. flame retardant in plastics, little antimony has been recovered from waste plastics to date. Many countries will have to rely on recovering antimony from refractory ores and metallurgical residues. For the purpose of efficient and selective antimony recovery with minimal economic cost and environmental impact, researchers proposed various innovative pyrometallurgical, hydrometallurgical, and electrochemical processes. This paper reviews the state-of-the-art technologies for the antimony separation and recovery from refractory ores and metallurgical residues. The advantages and limitations of these technologies are discussed with respect to antimony recovery efficiency and the impurity level of the products. Furthermore, future perspectives regarding technological trends and industrial applications are discussed.
Hydrogen peroxide (H2O2) is a strong oxidizer that causes non-selective oxidation of sulfide minerals, and its influence on bismuth sulfide ores is not well-documented. In this study, H2O2 was proposed as an alternative bismuthinite depressant, and its effect on a Mo-Bi-containing ore was intensively investigated by batch flotation tests. Results showed that the addition of H2O2 significantly destabilized the froth phase, thus decreasing the solids and water recovery. The recovery of bismuth in molybdenum concentrate was dramatically decreased to 4.64% by H2O2 compared with that in the absence of H2O2 (i.e., 50.14%). The modified first-order kinetic model demonstrated that the flotation rate of molybdenite slightly declined after H2O2 addition, whereas that of bismuthinite was drastically reduced from 0.30 min−1 to 0.08 min−1 under the same condition. Simulation revealed that H2O2 affected the floatability of both molybdenite and bismuthinite but resulted in more detrimental effect to bismuthinite. Hence, H2O2 has the potential to act as an effective depressant in bismuth sulfide ore flotation.
The polymetallic Cu–Zn ore of the Rockliden massive sulphide deposit in the Skellefte District in north-central Sweden contains a number of deleterious elements in relevant concentrations. Of particular concern is the amount of antimony (Sb) reporting to the Cu–Pb concentrate. The aim of this study was to compare different model options to simulate the distribution of Sb minerals in a laboratory flotation test based on different degrees of details in the mineralogical information of the flotation feed. Experimental data obtained from four composites were used for the modelling and simulation. The following different simulation levels were run (sorted from least to highest level of detail of their mineralogical information): chemical assays, unsized bulk mineralogy, sized bulk mineralogy and particle information. It was shown that recoveries simulated based on bulk mineralogy are mostly within the error margin acceptable in the exploration stage of the Rockliden deposit. Unexpected high deviation in the simulation using particle information from the original recovery has been partly attributed to the fact that recovery of non-liberated particles cannot be modelled appropriately in the present version of the modelling and simulation software. It is expected that the implementation of full particle information in simulation will improve the Sb distribution model for the mineralogically complex Rockliden deposit.
Some data on the bismuth recovery technologies in the production of copper and the processing of copper-bismuth and tungsten-molybdenum concentrates are given.
The Rockliden massive sulphide Zn–Cu deposit contains minor amounts of Sb minerals. The Sb mineralogy is complex in terms of composition, micro textures and mineral associations. The main Sb minerals comprise tetrahedrite, bournonite, gudmundite and Sb–Pb sulphides such as meneghinite. The presence of these minerals is especially critical to the quality of the Cu–Pb concentrate. To study how they are distributed in a simplified flotation circuit and what controls their process behaviour Sb-rich drill core samples were selected from the Rockliden deposit and a standard laboratory flotation test was run on the composite samples. Scanning electron microscope-based automated mineralogy was used to measure the Sb mineralogy of the test products, and the particle tracking technique was applied to mass balance the different liberation classes to finally trace the distribution of liberated and locked Sb minerals. The mineralogical factors controlling the distribution of Sb minerals are mineral grain size, the degree of liberation, and associated minerals. Similarities in the distribution of specific particle types from the tested composites point towards systematics in the behaviour of particles and predictability of their distribution which is suggested to be used in a geometallurgical model of the deposit.
The flotation of stibnite with Pb²⁺ as an activator was examined with micro-flotation tests, inductively coupled plasma mass spectrometry (ICP-MS) experiments and density functional theory (DFT) calculations. It was found that at a pH of 6.5, the addition of Pb(NO3)2 notably improved the flotation efficiency of stibnite with butyl xanthate (BX). At a pH of 6.5, Pb²⁺ was the dominant species adsorbing at the stibnite surface. DFT calculations implied that Pb²⁺ could adsorb at five different sites on the stibnite surface. Furthermore, BX could adsorb onto the Pb sites at the Pb-activated surface and the Sb sites of the un-activated surface, while the adsorption of BX on the Pb-activated surface is more stable.
Partial density of states (PDOS) analysis revealed that the poor overlapping between the S 3p orbital of BX and the Sb 5s orbital accounts for the weak interaction between BX and the un-activated surface. For the Pb-activated surface, the Pb 6p orbital and the S 3p orbital of BX were efficiently overlapped with each other, resulting in a relatively stronger interaction between BX and the activated surface. The DFT simulation provides a deep insight into the role of Pb²⁺ during the activation flotation of stibnite.
The depressive action of chromate and dichromate salts on galena was investigated theoretically and experimentally. The potential-pH diagram for the system Pb-Cr-S-H20 at 25°C was constructed. It showed that lead chromate has a stable domain over a wide range. Considering the pseudostable character of chromate ions, the stable domain of lead chromate extends to reducing conditions. Thermodynamic and kinetic considerations lead to the following reactions as proposed mechanisms of depressive action of chromates or dichromates on galena:3PbS + 11 Cr04 + 16H == 3PbCr04 + 4Cr203 + 3S04 + 8H20 3PbS + 5HCr04 + 5H+ = 3PbCr04 + Cr203 + 3S0 + 5H20 These reactions explain experimental observations such as the shift in pH of buffered solutions containing galena in the presence of cluomates, the enhanced adsorption of cluomates on the surface of galena in the presence of dissolved oxygen, and the reduced chromate adsorption on the galena surface above pH 9. Infrared spectroscopic analysis of the surface products on galena treated with cluomate or dicluomate solutions confumed the formation of lead chromate. Although the formation of chromic oxide or chromic hydroxide was not confirmed because of insufficient sensitivity of the method used, the recent work by Matsuoka, et al., supports the proposed mechanism.
Many kinds of minor elements as well as major constituent elements as well as major constituent elements are usually contained in sulphide minerals such as chalcopyrite, galena, sphalerite and pyrite. Such minor elements as As, Sb and Bi exert harmful influences on mineral processing and refining. In this paper, the mineralogical characteristics of As, Sb and Bi, as minor elements in the sulphide minerals and Kuroko ores, are discussed on the basis of available data. The flotation behavior of some sulphide minerals bearing As, Sb and Bi was evaluated.
Flotation of antimony-bearing minerals in a complex sulphide ore from the Rakkejaur deposit has been studied in laboratory experiments. Microprobe equipment has been used to distinguish the various types of antimony mineral present in the flotation products. The results show that effective depression of the iron-rich antimony minerals can be obtained on flotation in a lime alkaline environment at pH 11.5 without increasing the losses of silver. These antimony minerals float more strongly at about pH 5.
The flotation behaviour of jamesonite was investigated with sodium diethyldithiocarbamate (DDTC), ethyl xanthate (KEX) and ammonium dibuyldithiophosphate (ADDP) as a collector respectively and lime and potassium cyanide as a regulator respectively. The flotation of jamesonite was also dependent on pulp potential. The potential-pH range for flotation of the jamesonite- collector system was established.
Laboratory flotation tests were conducted on several ores and bulk concentrates to evaluate the effectiveness of chemical oxidizing agents as selective depressants for arsenopyrite. These studies were intended to complement independent electrochemical studies of the flotation surface chemistry of arsenopyrite. The effectiveness of oxidizing agents in depressing arsenopyrite improved as the pH increased above 7. Strong oxidizing agents could depress previously floated arsenopyrite. The selective depression of arsenopyrite from bulk pyrite-arsenopyrite concentrates was achieved through the use of an appropriate oxidizing agent.
The separation of antimony mineral impurities from a complex sulphide ore was studied. The flotation properties of these minerals are not well known. Advanced X-ray spectrometry was used in the mineralogical analysis and the chemical environments in the pulp at the time of separation were determined. The ferrous mineral gudmundite was depressed in lime- and soda-containing, alkaline environments. The lead-antimony mineral boulangerite floated more readily, and the lime alkaline process environment seems to be the most favourable for depression of this mineral, whose behaviour resembles that of galena. The copper-lead-antimony mineral bournonite floats very easily and is difficult to separate in a process environment that is designed to float chalcopyrite.
The flotation properties of the mineral jamesonite, which occurs as an impurity in some chalcopyrite concentrates, were investigated and compared with those of stibnite. Jamesonite floats readily at natural pH without activation and with xanthate as the collector. The mineral can be depressed with dichromate at natural pH. Jamesonite, like stibnite, is capable of floating to some extent with frother alone.
A new type depressant, BK510, which developed by Beijing General Research Institute of Mining and Metallurgy for the molybdenite and bismuth flotation separation was introduced. The conditional test and industrial test on replacing sodium sulfide by BK510 were made in Shi-Zhu-Yuan deposit in hunan of China and achieved very good results. Satisfactory metallurgy and economy have been achieved at laboratory and in commercial operation. Closed circuit experiment index sign was: molybdenite-bismuth concentrate contained 6.10 per cent of Mo and 6.84 per cent of Bi. It was acquire that molybdenum concentrate assay was 47.90 per cent Mo, its recovery was 97.27 per cent, it contained 0.40 per cent Bi, Bismuth rough concentrate assay was 7.75 per cent Bi, its recovery was 99.28 per cent, it contained 0.19 per cent Mo. The comparable flotation results to sodium sulfide showed that the depressing effect of BK510 on bismuth was same to sodium sulfide, and with small quantity of agent. In dosage of BK510 was 50~100 g/t, but dosage of sodium sulfide was 800~1000 g/t. In addition, BK510 was cheap and low in toxicity. Molybdenite recovery and grade getting in plant was similar to at laboratory scale. So, it was a quality depressant of Mo-Bi separation, and could substitute sodium sulfide completely as the depressant. The pollution of sodium sulfide can be avoided, and the production environment can be improved.
Molybdenite ore at Bando Mining Company in South Korea contains chalcopyrite, galena and bismuthinite as impurities. For the reason of metallurgical procedures, high grade flotation concentrate of Molybdenite has been demanded, and impurities such as Cu, Pb, Bi and W must be extremely low, for example Cu<0. 3%, WO3<0. 03% and Pb< 0.05%.Flotation tests were carried out for the purpose of producing the concentrate that contains Cu <0. 1%, Bi<0. 1% and Pb<0. 05%, and that the grade of MoS2 was over 85%. As the results of a series of flotation tests, we obtained the most satisfactory results by adopting the hot water conditioning by using Na2SO3 only or in combination with KCN and Na2S. It was recognized that the depressant action of of Na2SO3 was very effective to bismuthinite and galena, when conditioning treatment was done at high water temperature, inspite of its low effect at room temperature.
The copper industry is witnessing great interest in the development and utilization of copper-arsenic deposits. While most plants tend to use traditional processing technologies, the depletion of conventional copper ores has created competition for designing and implementing new process alternatives for the treatment of copper-arsenic ores containing minerals such as enargite, luzonite and tennantite. However, the downstream processing of enargite-containing concentrates represents a major metallurgical challenge. An important consideration in all enargite processing options is the deportment and stabilization of arsenic in a benign form that meets current and anticipated environmental regulations. Smelters are subject to penalties on high arsenic-bearing copper concentrates. While reductive roasting can drive-off the arsenic from enargite concentrates to an acceptable level suitable for smelting (
Five tannin extracts from different trees were used to improve jamesonite flotation in the Changpo plant in Nandan, China. The flotation results demonstrated that the larch tannin extract was the best for improving jamesonite flotation. Compared to flotation without larch tannin extract, industrial trial results indicated that the Sb and Pb grade in the final concentrate increased from 13.1% and 14.8% to 17.6% and 19.9%, respectively, as the Zn grade decreased from 9.1 to 6.5%. In addition, the Sb and Pb recovery of 84.5% was achieved with the larch tannin extract (84.2% without larch tannin extract). Larch tannin extract depressed not only marmatite, but also to some extent pyrrhotite, arsenopyrite, pyrite, calcite and quartz.
In the Somincor Neves Corvo ore, antimony occurs in the tetrahedrite-tennantite series of minerals while copper occurs in chalcopyrite and in other minerals present in minor quantities including the tetrahedrite-tennantite series. Current processing strategies do not directly address the control of antimony contamination of the final copper concentrate. The focus of this paper is to report initial findings of a major investigation undertaken on both a plant and laboratory scale.It was found that both collector type and concentration significantly affected copper/antimony selectivity. In particular, staged addition of dithiophosphate collector was required to control tetrahedrite flotation during copper roughing. However, size by size analysis showed that tetrahedrite flotation increased during scavenging due to further collector addition which was required to recover coarser chalcopyrite bearing particles. The production of separate rougher and scavenger concentrates enriched in chalcopyrite and tetrahedrite respectively, became an attractive option.It was found that oxidative treatment was required to adequately control tetrahedrite flotation. This treatment was applied to a final, low volume concentrate stream containing large quantities of tetrahedrite.In order to further enhance selectivity, fundamental work is aimed at establishing the relative ease of oxidation of chalcopyrite and tetrahedrite as well as the influence of oxidation: on the stability of the adsorbed hydrophobic species. A fiowsheet embracing these various concepts is proposed.
In order for the high-arsenic regions of the Northparkes copper–gold orebody to be beneficiated economically, tennantite ((Cu,Fe)12As4S13) present in the ore needs to be rejected to enable copper concentrates to meet the typical smelter penalty level of 2000 ppm As.Using a composite sample of high-arsenic drill cores from Northparkes it was possible to selectively separate tennantite from chalcopyrite (CuFeS2) and bornite (Cu5FeS4) using controlled-potential flotation. The separation was made on a bulk copper–arsenic concentrate after reducing the pulp potential to about −150 mV SHE at pH 12 and floating the tennantite from the other copper minerals. The basis of the separation relies on findings that the lower limiting pulp potential threshold for tennantite is lower than that for chalcopyrite such that there is a potential window in the reducing region where tennantite is strongly floatable but chalcopyrite is not. Little or no selectivity between tennantite and chalcopyrite was found in the oxidising pulp potential region for the range examined.From the composite sample tested, which had a head grade of 0.11% As and 1.2% Cu, it was possible to produce a low-arsenic high-copper concentrate containing 52% of the non-tennantite copper and assaying 2600 ppm As. Computer simulations have shown that for a feed containing a more typical arsenic and copper level (200 ppm As and 1% Cu) the efficiency of separation should be sufficient to concentrate about 61% of the copper in a product assaying less than 2000 ppm As.A conceptual flowsheet for arsenic rejection from Northparkes copper–gold ore, based on the findings from this study, is presented and discussed in this paper.
In an extensive programme of batch flotation tests on mixtures of purified minerals, it was established that freshly ground chalcopyrite displayed natural flotability in an oxidising environment and non-flotability in a reducing environment. No rational hypothesis to account for this behaviour has emerged. Grinding in an iron mill produced strongly reducing conditions and consequently suppressed flotation which was restored subsequently by raising the potential of the pulp either by aeration or by the addition of oxidants. The coarse particle sizes recovered more slowly than other fractions. The type and addition of frother had a pronouced effect on the natural flotability, but no proven effect on hydrophobicity. There is some evidence that whilst these observations apply to chalcopyrite from several sources when floated from mixtures with quartz, chalcopyrite in real ore samples does not necessarily show the same flotation behaviour.
Recent advances in fine particle processing of antimony oxide minerals are briefly reviewed. The physico-chemical characteristics of surfactant emulsion droplets, such as droplet size, interfacial potential and adsorption, are correlated with the floatability of fine antimony oxides in search of a new way to recover the fine oxides effectively. Fine antimony oxides were found to exhibit high floatability with ultrasonically prepared emulsions of dual surfactants (octylhydroxamic acid and Syntex) and hydrocarbon oil. The stability of the emulsion droplets is delineated by a mixed-film model of dual surfactant co-adsorption at the oil-aqueous solution interface. The mechanism of emulsion flotation is discussed.
The case of flotation of a ‘bulk’ hydrophobic mineral in the absence of a collector is considered. The theory developed takes into consideration the van der Waals and electrostatic forces and leads to a simple quantitative criterion of floatability containing the zeta-potential.Experiments with the flotation of antimonite confirm the theory and the quantitative form of criterion of floatability revealing the role of zeta-potential.
On the basis of knowledge of the beneficiation properties of antimony minerals reported in the previous part of this article, this second half reviews the state of technology for separation of those minerals and summarizes experience from processing plants. The high density of antimony minerals and their tendency to grind to slime (hardness of 2.5 on Mohs' scale) make gravity separation in the mill circuit an interesting possibility for the first step in the process. Gravity concentration is already found in some process layouts, but there are undoubtedly more applications where modern gravity separation equipment could be used. Marketing considerations make separation of arsenopyrite an important part of the process in some cases. Special cleaning processes have been developed for the purpose, but more attention needs to be paid to selectivity in the primary Sb flotation. Because of the rise in gold prices, some antimony ores should really be viewed as gold ores with antimony as a by-product. In this context, cyanide leaching is an obvious step in addition to gravity separation, and the most logical procedure here would be first to separate the antimony at natural pH and then to leach out the gold with cyanide. The design of flotation circuits for beneficiation of stibnite is usually very straightforward: rougher, scavenger and two-stage cleaner flotation are usually enough to produce concentrates grading better than 60% Sb. Recycling circulating returns to the middle of the rougher flotation circuit offers an attractive way of obtaining high-grade concentrates while achieving high capacities. The choice of methods for depressing antimony minerals in complex sulphide ores depends on what kind of mineral is the predominant impurity. It ought to be possible to depress ferrous antimony minerals effectively in pyrite-selective environments.
Antimony has long been reckoned as a strategic metal. It is used today not only in metallic form, e.g. as an alloying element in lead, but also to an increasing extent in the form of Sb2O3 as a flame retardant filler in plastics.The mineral raw material that provides nearly all the world's supply of antimony is stibnite, Sb2S3. The flotation properties of this mineral are fairly well established. Its crystal structure makes its floatability depend on which surfaces of the mineral are exposed by grinding. Factors of pulp chemistry, however — especially the pH of the pulp — seem to have a more decisive influence on the result of flotation. At about pH 4–5, stibnite floats spontaneously with acceptable recovery, and with collectors it floats very well. Activation with metal salt is needed at neutral pH, while floatability is seriously impaired at pH 10.The beneficiation properties of other antimony minerals have been very little studied. Antimony oxides, pure antimony and jamesonite, Pb4FeSb6S14, are also recovered as values in addition to stibnite. Antimony oxides respond very poorly to flotation with xanthate collectors. The flotation properties of jamesonite are more akin to galena than to stibnite, and this mineral floats best at around natural pH without activation.The high antimony content of some very complex sulphide ores is a problem in beneficiation, giving concentrates heavily contaminated with antimony minerals which, if they cannot be separated, increase the cost of smelting.These impurities can be grouped according to their beneficiation properties:View Within ArticleThe flotation properties of the lead-antimony minerals grow more like those of galena the higher they grade in lead. Bournonite, a fairly common mineral in complex sulphide ores, floats well in process environments where chalcopyrite normally floats; in many cases this makes it impossible to separate from chalcopyrite. Initial studied have indicated that the ferrous antimony minerals have properties resembling those of pyrite and pyrrhotite.
A study of floatability characteristics and surface phenomena of antimonite in solution of different flotation reagents at various pH values is aimed at determining the relationship between the floatability and the compostition of the antimonite surface layer.The investigation was carried out in the presence of depressing (a combination of FeSO4 and Na2S or NaCN) and activating (Pb-acetate) agents, with xanthate collectors (KEX and KAX), as a function of pH.A good correlation between the floatability and infrared spectrophotometric data was obtained. In conditions of unactivated antimonite, using KAX as a collector, Sb(AX)3 was suggested as the species responsible for hydrophobization of the mineral surface. In an alkaline medium, IR spectra do not indicate its presence on mineral surface. After activation by Pb2+ ions, Pb(AX)2 and Sb(AX)3 or only Pb(AX)2, depending on pH, were found in the surface layer of antimonite.No adsorption of KEX on unactivated antimonite surface was identified. When Pb-acetate was added, the IR spectra demonstrated the presence of Pb(EX)2 in the antimonite surface layer and complete floatability of the mineral was obtained.The Sb(AX)3 absorption bands were observed in the surface layer spectra of depressed antimonite treated with KAX in an acid medium, when good depression was not achieved. The full depressing effect was observed in alkaline medium and the spectra of surface layer showed no adsorption of amylxanthate.
Arsenic is a penalty element in base metal gravity and flotation concentrates and during beneficiation efforts are often made to reduce its level in concentrates destined for smelting. In some Australian tin and tantalum circuits this arsenic can occur as elemental arsenic, lollingite, or arsenopyrite. A great deal is known about the flotation of arsenopyrite, but little is known about the flotation of lollingite and arsenic. This paper is concerned with the flotation of metallic arsenic with ethylxanthate as a function of pulp pH and pulp potential (Eh). A synthetic mixture of arsenic metal and quartz was used in all tests.Arsenic was found to be strongly floating (up to 95% recovery in 8 min flotation with Aerofroth 65 frother and 40 g/t of KEX) over the pH range 5–10. At more alkaline conditions, the recovery dropped off slowly with increasing pH.At pH 6, arsenic was strongly floating over the pulp potential range + 125 mV to + 275 mV vs. SHE (Standard Hydrogen Electrode) but exhibited an upper limiting threshold value of about + 375 mV vs. SHE. The flotation response dropped off slowly with more reducing conditions below about + 125 mV vs. SHE. Here the flotation kinetics were slow. At pH 10, arsenic was found to be strongly floating in the potential range − 300 mV vs. SHE up to about + 225 mV vs. SHE. Interestingly, no lower limiting potential in the reducing potential range tested (down to − 300 mV vs. SHE) was identified. The rate data indicated fast flotation kinetics at pH 10 at potentials less than about + 225 mV vs. SHE.At pH 6, little genuine flotation of metallic arsenic in the absence of collector was observed and it appears that metallic arsenic does not exhibit any significant natural flotability as do some other metals.Importantly, the results of this study show that metallic arsenic could potentially be readily removed from base metal concentrates by controlled potential flotation over a wide range of pH values by a simple reagent combination.
Arsenic is one of the most dangerous inorganic pollutants and thus a penalty element in many base metal concentrates. Arsenic removal in sulphide flotation has been studied extensively with various approaches, including pre-oxidation of flotation pulp, Eh control during flotation and the use of selective depressants/collectors. Pre-oxidation of flotation pulp using oxidizing agents or aeration conditioning represents a simple approach in arsenic removal and was found effective in many cases. Selective flotation of arsenic minerals through Eh control has made significant advances in recent years with promising results achieved. In addition, various depressants and collectors have also been studied in arsenic removal. In this communication, the various approaches used in selective flotation of arsenic minerals are reviewed with emphasis on the development in recent years.
Voltammetric studies, contact angle measurements, collector and collectorless microflotation tests were carried out in this study to investigate the oxidation properties and flotation characteristics of enargite as well as chalcopyrite. Selective flotation of enargite from chalcopyrite under varied pulp potentials was conducted to investigate the feasibility of enargite removal from a chalcopyrite concentrate.The test results indicate that chalcopyrite began to oxidize quickly at a much lower potential than enargite. Enargite could be floated well at a potential higher than +0.2 V vs. SCE while chalcopyrite was completely depressed at a potential higher than +0.2 V vs. SCE. Selective flotation revealed that enargite can be successfully removed from chalcopyrite through controlling the pulp potential higher than +0.2 V and lower than +0.55 V vs. SCE.
Selective oxidation of minerals was investigated as a means to separate by flotation the copper sulfide minerals of chalcocite, covellite and chalcopyrite from the arsenic copper sulfide minerals of enargite and tennantite in mixed mineral systems. It was found that a separation of these minerals could be feasible after selective oxidation of their surfaces in slightly acidic pH conditions, or after oxidation and selective dissolution of the surface oxidation products with a complexant in basic pH conditions.
The floatability of enargite (3Cu2S.As2S5) has been determined as a function of pulp potential to establish whether the flotation behaviour of the mineral differs sufficiently from that of other copper minerals thus offering the prospect of rejecting arsenic from the Tampakan ore by potential control during flotation.The results of a single mineral flotation study reveal that a threshold potential exists above which enargite floats and below which it does not. On the standard hydrogen electrode (SHE) scale this potential is about − 75 mV SHE at pH 8 and about − 25 mV SHE at pH 11. By contrast, chalcopyrite (CuFeS2) has a threshold potential of about + 160 mV at pH 8 and chalcocite (Cu2S) has a threshold potential of about − 155 mV SHE at pH 11. In addition, chalcocite and cuprite (Cu2O), but not enargite and chalcopyrite, have an upper limiting potential at pH 11 (> + 270 mV SHE) above which their floatabilities are depressed.On the basis of these results it was concluded that the selective separation of enargite from other copper minerals would be enabled by controlling the flotation pulp potential. Specific flotation separations are as follows.•Copper sulphides such as chalcocite and copper oxides such as cuprite could be floated from enargite at pH 11 after setting the potential to about − 125 mV SHE.•Enargite could be floated from copper–iron sulphides such as chalcopyrite at pH 8 or pH 11 after setting the potential to about 0 mV SHE.•Enargite could be floated from copper sulphides such as chalcocite and from copper oxides such as cuprite at pH 11 after setting the potential to about + 290 mV SHE.A conceptual flow sheet for arsenic rejection at Tampakan, based on these findings, is presented and discussed in this paper.
Advances in interfacial phenomena of particulate/solution/gas systems. Applications to flotation research
- S Chander
- J M Wie
- D W Fuerstenau
Advances in interfacial phenomena of particulate/solution/gas systems. Applications to flotation research
- N P Finklestein
- S A Allison
- V M Lovell
- B V Stewart
Removal of Cu, Pb and Bi in molybdenite ore by flotation. Fusen, The Flotation Research Association of Japan
- H Matsubara
- T Ishihara
Flotation of unoxidized and oxidized sulphide minerals - antimonite, arsenopyrite, covellite, lollongite, marcasites, orpiment, pyrrhotite and tetrahedrite
- E Plante
Mineral processing of low-grade bismuth-tungsten ore at Akagane mine
- K Shoji
- Y Takamura
- A Kuroda
An examination of bismuthine activation by copper sulphate
- I T Tsvetkov
- S I Mitrofanov
- N D Mogilev
Sunshine mine: responding to mineral changes in ore feed
- J L Allen
Australasian Institute of Mining and Metallurgy
- L A Newcombe
Recent trends in the processing of enargite concentrates
- M S Safarzadeh
- M S Moats
- J D Miller
The use of pulp potential control to separate copper and arsenic - an overview based on selected case studies
- L K Smith
- K J Davey
- W J Bruckard
The case for the chemical theory of flotation
- A F Taggart
- Grm Del Giudice
- O A Ziehl
Plant practice: Sulfide minerals and precious metals
- J T Woodcock
- G J Sparrow
- W J Bruckard
- N W Johnson
- R Dunne
Proceedings of the XXI International Mineral Processing Congress
- Yen Wt
- J Tajadod