OpenUCT :: Browsing by Author "Corin, Kirsten"

Browsing by Author "Corin, Kirsten"

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Open Access
A stepwise Study on the characterisation and processing of South African Platinum Group Tailings
(2023) Manenzhe, Re?oket?we; Corin, Kirsten
The endurance of the mining industry has led to the near-depletion of some of the most processible ore types. This has resulted in a unique challenge that necessitates unique innovation for the industry. Firstly, existing technologies are increasingly geared towards improved efficiency in processing lower grade ores. Secondly, ores that were previously processed with older, and in some cases, inefficient technologies, have emerged as potentially viable solutions that would help maintain concentrator capacities. On the other hand, as the industry has endured, so has its waste accumulation. Tailings dumps continue to grow, and continue to pose various environmental issues. But although they are a waste product, tailings also have several merits to them. They are already mined and readily available, and reprocessing them immediately addresses two conundrums i.e. how can the industry source alternative ores, and how can it deal with the steady accumulation of waste? It is expected that as a result of their initial processing from so-called fresh ores, and their stay in their respective dumps, the tailings will be altered or tarnished. The surface properties of their compositional minerals will be oxidised and layered with other compounds that might hinder their interaction with flotation reagents such as collectors and hence, hinder their flotation response, presenting a challenge for their proposed reprocessing. More obviously, as they are a waste product, their grades will be far lower than the fresh ores that produced them. Numerous studies aim to elucidate the viability of the metallurgical reprocessing of tailings. However, this flurry of innovation thus far extends to the vast stretches of the Witwatersrand dumps, and thus, to gold. The issue with processing other tailings types then becomes threefold. Is there enough value in the dumps to justify tailings beneficiation? What beneficiation method would be suitable? Would the method yield economically viable results? This study acquired three bulks of PGM tailings to investigate these questions. The first bulk was Merensky, the second was UG2, and the third was a deslimed version of UG2 in which a portion of denser minerals were separated out. For the sake of convenience, the bulks are referred to by their shorthands throughout the thesis. Merensky became MER, the normal or raw UG2 became UG2R, and the deslimed UG2 became UG2D. The study conducted the investigations along two lines: first, by characterising the tailings, and secondly, by floating them. The first characterisation step was a full elemental and mineralogical analysis that quantified the amount of valuables as well as gangue minerals; this was done via XRD and QEMSCAN. The second step was to determine the degree of oxidation by using qualitative, in-situ experimental methods; these are EDTA extraction and oxygen reactivity. It was hypothesised that EDTA extraction would measure the concentration of secondary and oxidised materials on the mineral surfaces; and oxygen reactivity would measure the demand of oxygen in each tailings, and therefore the tailings' capacity to react with species in the pulp phase. The qualitative methods have only ever been used for fresh ores, and have shown to be reliable in predicting and/or explaining the flotation behaviour of those ores. They have never been used to predict an/or explain the flotation behaviour of an ore material that has already been processed, and is therefore very low grade, and oxidised. If they are as viable for tailings as they are for fresh ores, they would determine different EDTA extraction indices and oxygen demand constants. QEMSCAN and XRD provided the different concentrations of the different minerals across the tailings, and showed that some minerals were present in one tailings but not the other. For instance, MER contained the highest fraction of chalcopyrite, as well as the highest fraction of sulphides. UG2R and UG2D each had more than ten times the fraction of chromite seen in MER. MER, on the other hand, had more pyroxenes, plagioclase and amphibole. It was expected that these differences in composition would be the cause of the different extraction indices and oxygen demand constants. The robustness of both methods thus had to be tested. This was done by altering each of the tailings and testing whether the extraction indices and oxygen demand constants would change. The surface and composition alteration was induced with ultrasonication and desliming. The study thus answered the question: can the EDTA extraction and oxygen reactivity methods detect when a change has occurred on the mineral surfaces of the same tailings? The tailings were then floated, and it was here that the question of economic viability was assessed. For the concentrators from which MER, UG2R and UG2D were collected, a reprocessing venture would be deemed economically viable if 30% of the copper present in the tailings was recovered. The flotation performance was thus analysed with copper recovery as the primary positive indicator. Nickel behaviour was also tracked in case of any supporting elucidations, and also because pentlandite is the primary PGE-carrier for Merensky and UG2. The tailings were floated with DOW200 as the frother, SIBX as the collector, and CMC as the depressant. The results showed that the presence of the depressant resulted in very low solid quantities being recovered to the concentrate. In fact, less than 2% of each tailings was recovered, and less than 10% of the present copper (and therefore chalcopyrite) was extracted. When the depressant was removed from the reagent scheme, recovery of the solids improved to 10%, but the copper yield was still below 30%. So, the collector dosage was increased under the fundamental assumption that the hydrophobicity of the valuables would be improved, and in this way, the efficiency of separating the hydrophilic gangue from the valuables would also improve. The plan worked, and UG2R finally achieved the industrial objective of recovering 30% of the present copper. While MER and UG2D failed to do the same, their performance was also at its best under these conditions, with each tailings yielding roughly 23% copper recovery. In an effort to improve floatation, the tailings were cleaned via ultrasonication and desliming. These cleaning methods both had a detrimental effect on copper recovery. However, nickel (and therefore pentlandite) behaviour improved, showing that while the methods were disadvantageous to one mineral, they were favourable to another, and they might be useful for a study that uses a different metal as its positive performance indicator. The study also showed that MER, UG2R and UG2D have different copper EDTA extraction indices. UG2D had the highest index, followed by UG2R, and then MER. The copper minerals associated with UG2D can therefore be concluded to be more oxidised than those associated with UG2R and MER. Moreover, UG2D was the least reactive to oxygen, having an oxygen consumption rate constant of 0.113 min-1 when compared to 0.198 and 0.152 min-1 for UG2R and MER, respectively. Ultrasonication and desliming decreased each of these constants, indicating that when cleaned with the chosen methods, the mineral surfaces became less reactive to oxygen. And so, it was concluded that of the investigated tailings, UG2D was more oxidised than the other two, reacted the least with oxygen, and yielded the lowest copper recoveries. When UG2R achieved the highest reactivity with oxygen, it also yielded the lowest copper extraction index and the highest copper recovery. Overall, nickel behaved contrary to copper.
Open Access
An electrochemical investigation on the mechanisms of interfacial interactions of a xanthate collector on PGM surfaces in the presence of ions
(2023) Dzinza, Lucia; Corin, Kirsten; Tadie Margreth
Water is a vital transport and process medium used in mineral processing. Fresh water is substantially utilized as an ideal flotation media in the froth flotation process (Rao et al., 2016). However, the mining sector is impelled to save on the consumption of fresh water and reduce waste discharge owing to limited freshwater supply (Ridoutt and Pfister, 2010), stringent environmental regulations (Amezaga et al., 2011) and the increase in water demand among multiple users (Rijsberman, 2006). To improve water efficiency, the use of impure primary water supplies, and process water recycling has been implemented in flotation circuits. Generally, the recycling of process water is executed from tailings dams, thickener overflows, and dewatering and filtration units. The recycling of process water has been considered to be advantageous as it reduces freshwater consumption, lowers waste discharge and reduces volumes of reagents required in flotation circuits (Muzenda, 2010). However, recycled process water has been found to exhibit increased concentrations of typical contaminants such as colloidal matter, metal ions, sulphates, sulphites, thiosalts, calcium, magnesium, sodium, potassium, and residual floatation reagents. These contaminants affect the process water quality, which plays a vital role in flotation efficiency. Though a significant amount of work has been done on the effects of water quality on the flotation of valuable minerals, many studies have focussed on base metal sulphides and the use of benchscale techniques. Literature speculates that ions in process water hinder the interaction between xanthate collectors and valuable minerals, hence, contributing to a decrease in flotation recoveries (Kirjavainen et al., 2002, Boujounoui et al., 2015). The findings in literature have been deduced without an understanding of the underlying mechanisms of interaction between xanthate collectors and mineral surfaces in the presence of ions. Accordingly, literature still holds a lack of understanding on how ions affect the adsorption of xanthate collectors on mineral surfaces. This study, therefore, seeks to unpack the underlying mechanisms of interfacial interactions between ions with PGMs and the subsequent adsorption kinetics of a xanthate collector. This study investigated the effects of Ca2+, Mg2+, SO4 2- , S2O3 2- and Na+ ions at increasing ionic strength, on the adsorption of SIBX on synthetic PdS and PdTe2 minerals. The selection of the minerals was based on the need to give an insight into the differences in reactivities of the very floatable minerals (PdS) and the difficult-to-float minerals (PdTe2), with SIBX in the presence of ions. The mechanisms in question were examined by electrochemical techniques at laboratory scale. Rest potential measurements were used to determine the interactions of ions and/or SIBX on the PGM surfaces. Cyclic voltammetry was employed to determine the redox reactions that occur on the PGM surfaces in the absence and presence of ions and SIBX. Ultimately, electrochemical impedance spectroscopy was used to demonstrate the adsorption mechanisms of SIBX in the absence and presence of the investigated ions. The rest potential measurements generally displayed an increase in the extent of interactions between the investigated ions with the palladium minerals, with an increase in ionic strength. An inverse relationship was exhibited on the extent of interactions between the ions and PdS, and the extent of interaction between SIBX and PdS. Divalent ions displayed higher interactions with the palladium minerals than the monovalent ions investigated. All salts were found to demonstrate a decrease in the rest potential for PdS at all concentrations except for MgSO4, which increased the rest potential at 5 SPW and 10 SPW. Final rest potentials for most conditions were observed to be above the equilibrium potential of dixanthogen formation except for Na2S2O3 at 3 SPW, 5 SPW and 10 SPW, and CaCl2 at 1 SPW. Dixanthogen formation was most likely favoured on PdS for the conditions with final rest potentials above the equilibrium potential of dixanthogen formation. With regard to the PdTe2 mineral, it was found that most ions enhanced the interaction between SIBX and PdTe2. Contrary to the findings of PdS, it was found that most salts exhibited an increase in rest potential on PdTe2 except for Na2S2O3. Final rest potentials for all conditions investigated were observed to occur above the equilibrium potential of dixanthogen formation except for Na2S2O3 at all ionic strengths, MgCl2 at 10 SPW and NaCl at ionic strengths of 3 SPW, 5 SPW and 10 SPW. The latter conditions show that the formation of a metal-xanthate on PdTe2 was favoured. Generally, for both minerals, NaCl displayed the least interaction. It was found that increasing the ionic strength of salts, generally decreased the rate of dixanthogen formation on PdS. On the contrary, SIBX interacted more with PdTe2 at an increase in the ionic strength of salts. This observation favoured the formation of either a metal-xanthate or dixanthogen at a slower rate. Additionally, it was determined that the adsorption of ions investigated occurred via interfacial charge transfer kinetics, where an ion exchange mechanism has been proposed in the case of the divalent anions. In the case of divalent cations, it was presumed that the ions dissociate in solution and precipitate upon their interaction with the palladium minerals to hydroxides and/or carbonates. This study has shown that the mechanism of adsorption of ions on palladium minerals is heavily influenced by the type of mineral surface onto which the ions adsorb. The extent of interaction of ions with palladium minerals together with their corresponding oxidation products can be determined by the mineral type and the salt type and its ionic strength. Moreover, it was denoted that an electrochemical system that consists of salts at the palladium mineral surfaces can best be described by a resistor, Rs in series with a parallel circuit of a capacitor, Cdl, representing the electrical double layer and a resistor indicating Rct. For an electrochemical system with both salt and SIBX, it has been surmised that an equivalent circuit consisting of a resistor, Rs, in series with a parallel circuit of a capacitor, Cc, representing a coating layer formed on the palladium surfaces as a result of the adsorption and oxidation od SIBX and a capacitor, Cdl. This work has shown that the mechanisms of interactions between xanthates and PGMs in the absence and presence of salts can be successfully determined using electrochemical techniques. An understanding of such mechanisms developed from the interactions of Ca2+, Mg2+, SO4 2- , S2O3 2- and Na+ ions with SIBX on PGM minerals will help alleviate flotation problems caused by the troublesome ions. An understanding of the mechanisms proposed by this study will act as a diagnostic tool for developing flotation strategies that will maximize flotation recoveries where water quality is concerned
Open Access
An investigation into the effect of potential modifiers on the flotation of a copper sulphide ore
(2018) Dzinza, Lucia; Corin, Kirsten; Wiese, Jennifer
Oxidation, adsorption and reduction reactions are electrochemical in nature in the flotation of sulphide minerals which have semiconducting properties. Electrochemical mechanisms have two valuable implications in flotation, the potential across the mineral/solution interface determines flotation recovery and the anodic oxidation reaction involving the collector is an important parameter in imparting floatability. The reactions are dependent on the redox conditions in the pulp phase. Chemical control of redox potential (Eh) using potential modifiers may be exploited in flotation processes of sulphide minerals to improve their floatability, recoveries and grades; owing to the formation of a hydrophobic dithiolate or metal thiolate in the case of thiol collectors. In addition, chemical control of Eh is advantageous as it renders a more uniform electrochemical environment around the sulphide particles as compared to the external control of pulp potential. The adjustment of pulp potential using potential modifiers is being exploited as one of the main control parameters in sulphide flotation studies as it provides a diagnostic tool to develop flotation strategies and alleviate flotation challenges. Though potential modifiers have been previously investigated, no literature has addressed the correlation between their flotation performances on copper sulphides to their respective rest potentials at different concentrations. The present study explored the use of potential modifiers such as sodium hypochlorite (NaClO), potassium permanganate (KMnO₄) and potassium dichromate (K₂Cr₂O₇) on the flotation of a copper sulphide ore from Kansanshi Copper mine in Zambia. The potential modifiers were investigated at 1x10⁻⁴, 1x10⁻³ and 1x10⁻² mols which gave rise to various Eh values for each modifier. Batch flotation and froth stability tests were carried out at the ore’s natural pH whilst varying Eh. The dynamic stability factor (Σ) was used to quantify froth stability. Electrochemical techniques have been considered as an appropriate approach in the study of collector-mineral interactions. To complement results obtained from batch flotation and froth stability tests, rest potential measurements were carried out to determine the characteristic species formed on the chalcopyrite mineral surface at specific conditions. The potential modifier-collector-mineral interactions were investigated through rest potential measurements using the aforementioned potential modifiers, a thiol collector sodium iso-butyl xanthate (SIBX) and a pure chalcopyrite mineral. It was hypothesized that assuming an X⁻/X₂ equilibrium potential below 100 mV for SIBX, a redox potential range of 100-400 mV promotes good copper floatability due to the formation of dixanthogen and thus hydrophobic mineral particles which would result in a moderately stable froth. Rest potentials above 500 mV were hypothesized to reduce copper floatability due to the presence of very hydrophobic mineral particles, which would increase bubble coalescence and bubble breakage or result in highly stable froth. In this study, the equilibrium potential of SIBX at 6.24x10⁻⁴ M was measured to be 80 mV. Furthermore, equilibrium potential of SIBX was determined to be concentration dependent. Rest potential measurements for all conditions investigated were in excess of the measured equilibrium potential, therefore implying that the dixanthogen species was formed as postulated. It was found that an increase in concentration of potential modifiers increased froth stability or bubble coalescence depending on the potential modifier used. Furthermore, concentrations of potential modifiers resulting in Eh values of 137-476 mV resulted in high copper recoveries >88%, with 1x10⁻² mols of KMnO₄ at 540 mV giving a very low copper recovery of 4.8%. However, though high copper recoveries were obtained between concentrations that gave rise to an Eh range of 137-476 mV, a slight decrease in copper recoveries of approximately <4%, was observed with even larger increases in concentrations of potential modifiers. The findings of this study showed that the use of potential modifiers improved copper grades as a result of the reduction in gangue material recovery. In addition, the present study has shown that though concentration or Eh induced by potential modifiers may affect the flotation performance of sulphide ores, the most dominant factor that has shown to have a greater impact is the nature of the potential modifier. Comparing the findings of this work to literature findings for NaClO, it was determined that different sulphide minerals indeed exhibit different rates of redox reactions at given conditions. Ultimately, an inverse relationship was determined to exist between copper recoveries and rest potential measurements. This study has provided insight into the use of potential modifiers in the flotation of copper sulphides from an electrochemical perspective.
Open Access
An investigation into the effects of pulp chemistry under wet and dry grinding on the flotation response of pyrite
(2018) Tseka, Relebohile; Corin, Kirsten; Wiese, Jennifer
Considering the depletion of high-grade ore deposits, the mining industry is faced with the challenge of processing low grade and more complex ores in order to meet the growing demand for metals and metal products. Therefore, it is of paramount importance to have a fundamental understanding of minerals processing operations in order to improve the recoveries of valuable metals on an industrial scale. It has been acknowledged that the chemical conditions during grinding as well as pulp chemistry have a significant influence on the recovery and selectivity of most sulphide minerals in the flotation process. Floatability of ores is mostly determined by surface properties and the surface properties are essentially controlled by the grinding conditions. The flotation response of sulphide minerals is influenced by factors such as: collector-mineral interactions, mineral surface oxidation, deposition of iron hydroxides/oxides from grinding media and the attachment of inorganic ions on the surfaces of minerals. These factors are on the other hand affected by dissolved oxygen (DO), pH, ionic strength of process water and other pulp chemistry factors. With the highly instrumented Magotteaux Mill® , the effects of these variables may be investigated during grinding. Several studies have shown that the grinding environment plays a vital role in the selectivity and recovery of sulphide minerals. During wet grinding, water allows the flow of electrons within the pulp (galvanic interactions between minerals themselves and minerals and grinding media). Pyrite is reactive and can easily oxidise when exposed to air or oxygen. Pyrite and most sulphide minerals are more inert than the electrochemically reactive grinding media. Therefore, during grinding, grinding media come into frequent contact with sulphide minerals and a galvanic couple is created between the grinding media and sulphides. Due to galvanic interactions, oxygen reduction occurs on the sulphide mineral surface and iron oxidation takes place on the steel media. The redox reaction results in the formation of iron oxy-hydroxides on the surface of sulphide minerals. The oxy-hydroxide species prevent the adsorption of collector onto the mineral surface, making the mineral less floatable. Dry grinding limits the galvanic interactions present during wet grinding, due to the absence of water. Studies have been conducted and it has been shown that dry grinding yielded significantly less media wear relative to wet grinding owing to the absence of corrosive abrasion in the form of electrochemical oxidation of media during grinding. Reduced grinding media wear may imply that lesser iron hydroxide precipitates build up on the surface of the mineral hence improving collector adsorption and subsequently mineral recovery. Therefore, this suggests that dry grinding could result in improved sulphide mineral recovery as compared to wet grinding. It is necessary to consider the fundamental aspects of both grinding and flotation in order to improve concentrator performance as well as sulphide mineral recovery in the presence of nonsulphide minerals. Previous studies have investigated the influence of the grinding pulp chemistry factors on the flotation response of pyrite and other pure sulphide minerals. The possible influence that the presence of a non-sulphide gangue mineral may have during grinding and flotation has been ignored. The non-sulphide gangue cleans the surface of the sulphide minerals. Studies have shown that presence of quartz influences the formation of layers of hydrophilic species on the surface of sulphide minerals. The metal hydroxides will preferably deposit on the surface of non-sulphide mineral such as quartz rather than sulphide minerals. These studies also did not investigate the combined effects of pulp chemistry factors under dry and wet grinding. It should be noted that it is not possible to control pulp chemistry during dry grinding, thus these variables are controlled in the flotation cell in order to understand their effect on mineral surface after dry milling on pyrite flotation recovery relative to how they change the minerals surface properties during grinding. Change in chemical, surface properties of sulphide minerals can take place during milling and froth flotation. Therefore, this study aims to investigate the effects of DO, pH and grinding media type (forged steel and ceramic media) during milling and flotation process on the flotation response of pyrite (sulphide mineral) in the presence of quartz (non-sulphide gangue material). Wet milling was conducted in a Magotteaux Mill® while a Sala Batch grinding mill was used to carry out dry grinding. DO concentration and pH were controlled and measured in situ during wet grinding and were manipulated inside the flotation cell after dry grinding. The effects of the DO and pH, with changing grinding media type, on water and solids recovery, pyrite recovery and grade as well as flotation kinetic constants were studied. The EDTA extraction technique was employed to quantify the percentage of extractable oxidized iron leached from the mill product. The findings of this study have shown that under both wet grinding and dry grinding, an increase in pH from 9 to 11 resulted in increased water and solids recovery due to an increase the total concentration of OH ions in the system which led to increased froth stability owing to the reduction in pulp bubble size, as well as reduced bubble coalescence. This shows that the control of pulp chemistry during milling and flotation affected flotation process in the same way. The study has further shown that the highest recovery of pyrite, 100%, was achieved with inert grinding media (ceramic) under dry grinding. This might be due to cleaner pyrite surfaces created during dry grinding, since the prevention of media corrosion may lead to improved recoveries. During wet grinding, iron hydroxide is generated and reduces the flotation response of pyrite. Dry grinding generally produces much faster pyrite flotation kinetics than wet grinding because of the generation of particles with high surface energy and that leads to highly activated particles. It was therefore concluded that the grinding environment indeed has an effect on the flotation response of pyrite in the presence of gangue. This study has shown that careful manipulation of pulp chemistry, selection of grinding media and grinding environment may be used to manage pyrite recoveries within flotation.
Open Access
Considering the Action of Degrading Water Quality on the Electrochemical Response of Sulphide Minerals
(2021) Ndamase, Nolihle; Corin, Kirsten; Tadie, Margreth
Mining operations in arid regions are compelled to reduce their consumption of fresh water. Closed water circuits are an attractive solution and, in addition to reducing freshwater consumption, they have the added benefit of reducing reagent consumption as well as the environmental impact of mining operations by eliminating effluent discharge. Water is used as a liquid medium as well as a means of transportation during mining operations. Flotation, in particular, is a water intensive process where water makes up about 80- 85% of the pulp phase. Process water contains organic and inorganic species which accumulate as they are recycled. To avoid the treatment costs of removing these contaminants, many mining operations allow the quality of their water to degrade over time. When this water is introduced into flotation circuits the pulp chemistry is altered. Ionic species that accumulate in recycled process water have been shown by previous studies to be especially deleterious to flotation performance. Such ions include Ca2+, Cu2+, Mg2+, Pb2+, SO4 2- and S2O3 2- , to name a few. One of the effected flotation sub-processes is collector adsorption which is responsible for inducing hydrophobicity on valuable mineral surfaces. Accumulating ionic species have been shown to hinder collector adsorption which reduces the recovery of valuable minerals to the concentrate. Consequently, degrading water quality threatens the economic viability of mining operations that make use of closed water circuits. Xanthates are the most widely used collectors to treat base metal sulphide minerals. Most sulphide minerals are semi-conductors and xanthate adsorption onto their surfaces takes place via electrochemical reactions. Numerous studies have investigated the effect of degrading water quality on xanthate adsorption however there is very limited understanding as to how this takes place from an electrochemical perspective. This study therefore aims to investigate how the presence of accumulating ionic species at varying concentrations affects the electrochemical adsorption of sodium ethyl xanthate onto chalcopyrite and galena. Synthetic plant water (SPW) was used to mimic the composition of recycled process water. The synthetic plant water used in this study comprised of single salt solutions to isolate the effects of the ionic species of interest which were Cl- , Mg2+, SO4 2- and S2O3 2- . The salts used were NaCl, MgCl2, MgSO4, Na2SO4 and Na2S2O3. The ionic strengths were varied between 0.0242 M, 0.0727 M, 0.1212 M and 0.2426 M which correspond to 1, 3, 5 and 10SPW, respectively. Measuring the mineral rest potentials is an electrochemistry technique that was used to monitor how changing water quality affected xanthate-mineral interactions. Microflotation measurements were conducted to observe how changing water quality affected mineral floatability. Zeta potential measurements were used to determine the mineral surface charge and assess which of the ions of interest were active on the mineral surfaces. This study also aimed to investigate if electrochemistry techniques such as rest potential measurements can be used to predict flotation performance with changing water quality. Microflotation measurements revealed that degrading water quality only had an impact on the initial flotation kinetics of chalcopyrite and galena with both minerals achieving final recoveries greater than 90% regardless of changing water quality. The only exception was galena in the presence of the S2O3 2- ion where final recoveries were no greater than 35%. The S2O3 2- ion was found to be the most deleterious ionic species on both chalcopyrite and galena floatability, more especially on the latter. This was attributed to the formation of xanthyl and metal thiosulphate species which passivate the mineral surfaces, hindering collector adsorption. The formation of xanthyl thiosulphate species also reduces the amount of xanthate available for inducing hydrophobicity. It was hypothesized that increasing the ionic strength of the synthetic plant water would result in a greater hindrance of collector adsorption due to the competition between ionic species and xanthate for adsorption onto the mineral surface. This was partially true as the initial chalcopyrite recoveries in the presence of the S2O3 2- ion decreased with increasing ionic strength. Additionally, the highest initial galena recoveries in NaCl, MgCl2 and Na2S2O3 were achieved at 1SPW indicating that, even though the correlation was not linear, the initial galena recoveries in these salts decreased with increasing ionic strength. Contrastingly, the highest initial chalcopyrite recoveries in NaCl, MgCl2 and Na2SO4 were achieved at 10SPW indicating that increasing the ionic strength resulted in higher recoveries. This improvement in initial recovery with increasing ionic strength was attributed to the compression of the electrical double layer which causes the mineral-water and air-water interfaces to destabilize thus reducing bubble-particle attachment time. Rest potential measurements revealed that dixanthogen and lead xanthate were likely the dominant surface reaction species formed on chalcopyrite and galena during collector adsorption, respectively. The only exceptions were chalcopyrite in the presence of Cl- at 5SPW and S2O3 2- above 1SPW where the species that was favoured to form in these conditions was cuprous xanthate. Rest potential measurements also revealed that the presence of the S2O3 2- ion hindered xanthate-mineral interactions to a greater extent than any other ion of interest. It was hypothesized that rest potential measurements can be used as a quick and easy technique to assess the effect of changing water quality on the flotation performance of sulphide minerals. This is due to the ability of rest potential measurements to indicate the extent of xanthate-mineral interactions. Unfortunately, rest potential measurements failed to consistently predict flotation performance with changing water quality. They did however successfully predict the depressant effect of the S2O3 2- ion on the floatability of both minerals, especially that of galena. Zeta potential measurements indicated that all the ions of interest were active on the mineral surfaces. This confirmed the assertations made in previous studies that these ionic species are surface-active counter-ions that hinder collector adsorption and therefore mineral floatability due to mineral surface passivation. For example, the surface activity of the divalent cation Mg2+ had a stronger effect on mineral floatability than the monovalent Cl- anion. This was proposed to be due to the formation of the insoluble metal hydroxide Mg(OH)2 which rendered the mineral surfaces hydrophilic resulting in lower recoveries.
Open Access
Considering the action of frothers under degrading water quality
(2020) Tetlow, Sarah; Corin, Kirsten; Manono, Malibongwe
Froth flotation is a highly water-intensive process which is under scrutiny due to scarce fresh water supplies and increasingly strict environmental regulations with regards to polluted water discharge. This is driving the mining industry to use recycled water for their operations, which is usually sourced from tailings dams or concentrator thickeners. This means that the recycled water can contain elevated levels of dissolved solids which consist of various ions and other contaminants such as residual reagents. This presents a problem in the flotation circuit as these dissolved solids tend to affect the water quality and can impact the efficiency and performance of flotation operations. The stability of the froth is known to strongly affect flotation performance and thus the grade and recovery of the valuable minerals. Literature shows that both frothers and ions reduce bubble coalescence, and stabilise the bubbles that form, resulting in greater froth stability. Considering that the level of ions in process water is on the rise, and both variables act on the froth in a similar manner, it is becoming increasingly important to understand how frothers behave under conditions of increased ionic strength. If it can be determined how these variables interact, then it may be possible to manage frother dosage in operations that recycle process water with the aim of reducing the quantity and cost of frothers and limiting the need for large amounts of fresh water, while still maintaining flotation performance. Therefore, this study was undertaken to investigate how frother dosage and ionic strength, both individually and simultaneously, affect the froth stability and therefore flotation performance. This study was limited to varying the frother type, frother dosage and ionic strength whilst keeping all other experimental conditions constant. Batch flotation tests were carried out involving the bulk flotation of chalcopyrite and pentlandite. Flotation performance was evaluated by examining the water, solids, copper and nickel recoveries, and the grades of both copper and nickel. The ore used for this study was Kevitsa ore from Finland. Both the individual effects of frother dosage and ionic strength and their simultaneous action were analysed. It was found that increasing the frother dosage stabilised the froth and increased the recovery of water and solids but had no impact on the recovery of copper and only a slightly positive influence on the recovery of nickel. At the same time, the grades of both copper and nickel were found to decrease, likely due to increased gangue recoveries. Increasing the ionic strength also stabilised the froth which increased the recovery of water and solids, but both the recoveries and grades of copper and nickel were not significantly affected. Examining both variables simultaneously revealed that ionic strength was more influential than frother dosage in the recovery of water with the opposite being true for the solids recoveries. This means that a simultaneous increase in ionic strength and decrease in frother dosage by the same amount will increase the water recoveries and decrease the solids recoveries. It will also slightly decrease the nickel recoveries while having no effect on the copper recoveries. The grades of both will either increase or remain the same. Overall, managing the frother dosage under conditions of increased ionic strength, while still maintaining flotation performance, is possible and could result in a decrease in the quantity and cost of frothers required for flotation. It may also allow the mining industry to recycle more of their water without the need for extensive cleaning which in turn will reduce the amount of fresh water required for flotation and reduce environmental discharge. However, because ionic strength and frother dosage have varying levels of influence and therefore must be monitored, the amount by which the ionic strength of the water is allowed to increase, and the amount by which the frother dosage is decreased, need to be tailored to suit the needs of the plant with regards to water recovery and the recoveries and grades of the valuable minerals.
Open Access
Considering the effect of pulp chemistry during flotation on froth stability
(2016) Sheni, Nanji Ruth; Corin, Kirsten; Wiese, Jenny
On an industrial scale the need for improved flotation performance is of high importance in the current economic climate. Studies have shown that the pulp phase chemistry has a strong effect on the froth phase and therefore it is necessary to investigate how the manipulation of pulp chemistry factors can improve flotation performance. Research into the manipulation of this chemistry is well underway and factors including the pulp potential (Eh), pH, dissolved oxygen (DO) and ionic strength (IS) govern the pulp chemistry. This study aims to investigate how the manipulation of these factors affects the froth stability, bubble size and entrainment of the froth phase through Platinum Group Metal (PGM) flotation. In this study the Eh, pH, DO and IS were successfully manipulated to investigate their effects on froth stability and water recovery in 2-phase, as well as their effect on water and solids recovery, entrainment and the grades and recoveries of valuable minerals (copper, nickel, platinum and palladium) in 3-phase in the absence and presence of depressant at high dosages; 500 g/t Carboxymethyl Cellulose (CMC). Stability column tests were used to determine froth stability as a function of the dynamic stability factor (Barbian et al., 2005) and batch flotation tests were used to obtain the total water and solids recovered, the grades and recoveries of the valuable minerals as well as to determine entrainment. Further tests were performed to investigate the effect of changing the pH on the Eh in a 3-phase system in which all the other pulp factors were kept constant. The effect of changing the pulp factors on the froth bubble size was investigated by capturing side view images of the froth obtained in a batch flotation cell as each pulp factor was changed. This study has shown that careful control of the pulp chemistry, namely increasing IS, increasing pH, decreasing DO and decreasing Eh, resulted in improved froth stability. The Eh was found to be inversely proportional to the pH. This study has further shown that increased water recoveries and reduced bubble size in the froth were observed at 5 IS as compared to 1 IS due to the froth stabilising nature of the pulp at 5 IS. Operating at high Eh (500-730 mV) was observed to be detrimental to valuable mineral grades and recoveries and promotes entrainment. This kind of knowledge contribution may be key in improving flotation performance and increasing the grades and recoveries of valuable minerals obtained in South Africa's PGM mining industry.
Open Access
Editorial for Special Issue “Water within Minerals Processing”
(Multidisciplinary Digital Publishing Institute, 2022-03-14) Corin, Kirsten; Smart, Mariette; Manono, Malibongwe
The products of mining are key to the technology development of the future [...]
Open Access
Effect of VRM on a polymetallic sulfide ore and the flotation response as compared to conventional wet and dry rod milling
(2019) Nyakunuhwa, Hebert Simbarashe; Mainza, Aubrey; Corin, Kirsten
Comminution is an energy intensive, size reduction and mineral dressing process which consumes up to 50% of concentrator energy consumption. Conventional methods use mainly a combination of crushers and tumbling mills in comminution circuits. Energy consumption in these circuits has been found to be relatively high. To reduce the energy requirements, compression grinding equipment, Vertical Roller Mills (VRMs) and High-Pressure Grinding Rolls (HPGRs) have been identified as potential solutions, and they have been adopted in the cement industry. Reports from plants where these technologies have been installed in circuits indicate they are more energy efficient than the conventional comminution circuits. Studies have also suggested that the use of VRMs results in comminution products with relatively higher mineral liberation degrees. Unlike in the cement industry, comminution equipment in mineral processing circuits are also required to produce particles that can be separated and recovered in downstream processes. Froth flotation is a selective separation process that utilises differences in surface properties to separate value minerals from unwanted gangue. The success of flotation is dependent on chemistry, operational and equipment factors. The chemistry factors consider the interaction between flotation reagents and solids particles surface. The operational factors consider the effect of particle size distribution, mineralogy, feed rate, pulp density, pulp potential (Eh), bubble size, temperature and circuit design on flotation. The use of different comminution procedures may result in flotation feeds of different particle size distributions (PSDs), mineral liberation characteristics and pulp potential. Due to these differences, the resultant flotation response may differ. The present study was aimed at assessing the particle size distribution, mineral liberation profiles and the flotation response from material comminuted using the VRM floated under batch flotation conditions in a 3 litre Barker flotation cell. A complex polymetallic sulfide ore containing chalcopyrite (1.3 %), galena (2.4 %) and sphalerite (1.8 %) as the main value minerals and magnetite (68.0 %) and quartz (15.7 %) as dominant gangue minerals was used for the study. The ore was milled to target grinds of 55 %, 60 %, 65 %, 70 % and 75 % passing 75 µm respectively, at a grinding pressure of 600 kPa, air temperature of 300 K. For the benchmarking grind of 65 % passing 75 µm, the ore was also milled using heated air of temperature of 373 K and at elevated grinding pressures of 800 kPa and 1000 kPa. Further work was performed to evaluate if the VRM results are comparable to conventional dry and wet rod milling products floated under the same batch flotation conditions. An increase in grinding pressure was observed to result in an increase in throughput and a general decrease in specific energy consumption without a change in product particle size distribution nor the recovery of chalcopyrite, galena and sphalerite. Using heated air (373 K) resulted in the production of slightly less fines in the comminution products. The recovery of chalcopyrite, galena and sphalerite were not affected by the change in operating temperature. However, concentrate grade (selectivity) was compromised at elevated temperatures of comminution probably due to surface oxidation. The results indicated that the grind range to achieve the best flotation performance when using the VRM as a comminution device is between 60 % and 70 % passing 75 µm. The results also indicated that at the benchmarking grind of 65 % passing 75 µm, the specific energy consumption for comminution using the VRM was 54.3 % lower than that of the conventional tumbling mill circuit. The grind of 55 % passing 75 µm resulted in lower flotation efficiencies as the minerals were unlikely liberated enough whereas the grind of 75 % passing 75 µm resulted in poor performances due to low water recovery. Comparing VRM with wet and rod milling, the different comminution procedures resulted in flotation feed of similar PSDs for all grinds compared. The wet and dry rod milling products of grinds 55 % and 75 % passing 75 µm achieved better recoveries of chalcopyrite, galena and sphalerite as compared to the VRM performance mainly due to high water recoveries achieved. While mineral recoveries were above 90 % for the grinds of 60 % and 70 % passing 75 µm, the rod milling products had statistically better flotation recoveries at 95 % confidence compared to the VRM products. The mineral recoveries after dry rod milling were marginally better than after wet rod milling due to the minimisation of galvanic interactions during dry rod milling. For the benchmarking grind of 65 % passing 75 µm, VRM grinding resulted in 84 %, 84 % and 90 % liberated chalcopyrite, galena and sphalerite respectively. The liberation of chalcopyrite, galena and sphalerite after wet and dry rod milling were 80 %, 78 % and 90 % respectively. Chalcopyrite recovery was 96.7 %, 96.3 % and 96.7 % for the VRM, dry rod mill (RD) and wet rod mill (RW) products respectively. Galena recovery was 94.3 %, 94.3 % and 92.9 % for the VRM, RD and RW products respectively. Sphalerite recovery was 96.6 %, 97.4 % and 97.4 % for the VRM, RD and RW products respectively. The differences in recovery were statistically insignificant at 95 % confidence. Liberation differences did not translate to differences in recoveries as the ore was coarse grained. The recovery kinetics were very fast and independent of comminution procedure. Reference to the benchmarking grind therefore, the VRM can be retrofitted into existing plant installations as it is more energy efficient and the flotation performance was similar when using the flotation procedure tailored for tumbling mill-flotation systems.
Open Access
An electrochemical investigation of platinum group minerals
(2015) Tadie, Margreth; Corin, Kirsten; Wiese, Jenny
The Bushveld complex is the largest ore body in the world hosting platinum group elements (PGEs). It is a stratified orebody with three major reefs namely, the Merensky reef, UG2 reef and the Platreef. Platinum and palladium are the most abundant PGEs found in the Bushveld complex. They occur in the form of minerals/mineral phases with elements such as sulphur, tellurium, arsenic and iron. These minerals/mineral phases are associated with base metal sulphides occuring along grain boundaries. Unlike the Merensky and UG2 reef, the Platreef is almost barren of PGE sulphides and the distribution of base metals sulphides and their association with PGMs is erratic. Froth flotation targeted at the recovery of base metal sulphides is implemented in PGM concentrators to concentrate PGMs. Flotation of sulphide minerals is achieved with the use of thiol collectors to create hydrophobicity, and copper sulphate is often used to improve hydrophobicity and therefore recovery. Sodium ethyl xanthate (SEX) and sodium diethyl dithiophosphate (DTP) are commonly used as collectors on PGM concentrators. The erratic mineral variations in the Platreef ore, however, raise the question of the effectiveness of the application of sulphide mineral flotation techniques on this ore. Previous work by Shackleton, (2007) investigated the flotation of PGE tellurides, sulphides and arsenides. The study highlighted that the mechanisms with which these minerals interact with collectors and with copper sulphate was poorly understood. It is as a result of the findings of Shackleton's work that this study aims to elucidate the fundamental interactions of telluride and sulphide PGMs with thiol collectors and with copper sulphate. Subsequently this work also aims to compare the behaviour of these reagents on sulphide PGMs and telluride PGMs.
Open Access
Investigating electrolyte-reagent-mineral interactions in response to water quality challenges in the flotation of a PGM ore
(2019) Manono, Malibongwe Shadrach; Corin, Kirsten; Wiese, Jennifer
Froth flotation is a physicochemical process that enables the separation of valuable minerals and unwanted gangue minerals contained in an ore. It utilises the differences in surface properties of the minerals to be separated. Among other factors affecting flotation, water is a major factor as it acts as a reagent and transport medium. Therefore, it stands to reason that the quality of the water used in the process matters. Current environmental restrictions on water usage which are aimed at addressing the global scarcity of water require that mining operations recycle and reuse water within their operations. This necessitates proactive management strategies and initiatives aimed at understanding the impact that water could have on flotation and other water intensive processes. The development of such initiatives relies on the provision of sound and fundamental scientific evidence examining and decoupling the effects of water quality on the sub-processes of flotation. This would enable the creation of alternative operating conditions at which flotation could still occur without significant effects on production and profitability. Recycling of process water has for many years been the mining industry’s solution to reducing reliance on municipal water because mining operations are often located in arid regions. It has become clear that the recirculation of water in flotation circuits results in the accumulation of dissolved solids, electrolytes, unspent reagents and biological matter possibly resulting in poor flotation performance or alternatively high costs associated with water treatment. Given that water is both a reagent and transport medium in flotation, changes in its quality can significantly affect flotation performance through electrolyte-reagent-mineral interactions. This study has investigated whether there are any dominant or synergistic electrolyte-reagent interactions occurring during flotation which may impact negatively on the flotation performance. Interactions occurring in both the pulp phase and the froth phase were investigated through established bench scale flotation techniques. On the basis of available literature, investigations were carried out to identify inorganic electrolytes which had the biggest impact on froth stability as well as those which had a dominant role on depression, and specifically CMC efficacy on gangue management. A Merensky ore, typical of the South African Bushveld Igneous Complex was selected as previous work within the Centre for Minerals Research was conducted on ores of similar mineralogy. Threephase bench scale flotation and froth column tests were performed to examine the effect of increasing ionic strength of plant water and CMC dosage on froth stability and gangue recovery. Two-phase batch flotation and froth column tests were performed at various electrolytic conditions to assess the effect of ionic strength, electrolyte type and pH on froth stability using water recovery, foam height and foam collapse time as key performance indicators of froth stability. Settling tests, adsorption studies, zeta potential measurements as well as inorganic electrolyte speciation determination were considered in order to elucidate the role of water quality on gangue depression. Talc and pyrrhotite were selected as proxies for naturally floatable gangue (NFG) and sulphides respectively in order to simulate the possible behaviour of a Merensky ore. Increasing the ionic strength resulted in increased solids and water recoveries suggesting an enhancement in froth stability. When the effect of ionic strength on CMC behaviour was investigated under changing pH, results showed that contrary to findings at pH 9 which showed increases in solids recovery with increasing ionic strength, solids recovered decreased with increasing ionic strength at pH 11. This suggested that at higher pH levels above pH 10 there are hydroxy species present which inhibit the floatability of mineral particles either by forming layers on the mineral particles which hinder the action of the collector or through depressant efficacy enhancement. The speciation diagrams indicated that beyond pH 10, species such as CaOH+ increased in concentration especially at the higher ionic strength. Furthermore the zeta potential results for talc and pyrrhotite showed that at pH 11, the potentials were less negative compared to pH 9 for all synthetic plant waters proving that at pH 11 the pulp chemistry would exhibit a more depressive nature onto mineral particles owing to increased concentrations of positively charged hydroxo species at pH 11 compared to pH 9. These hydroxy species such as CaOH+ would adsorb onto the negatively charged mineral particle, reducing the negative surface charge of the mineral particle. Water recoveries increased with increasing ionic strength at both pH 9 and pH 11. These findings were further supported by 2-phase froth column tests in which water recoveries, foam height, and foam collapse time increased with increasing ionic strengths. This increase in froth stability with increasing ionic strength at both pH conditions is attributed to an increase in the [Ca2+], [Mg2+] and [SO4 2- ] which reduces bubble coalescence. Upon the determination of NFG recovery, entrained gangue recovery and total gangue recovery, it became clear that, at increasing ionic strength, there was a decrease in the recovery of NFG and entrained gangue per g of water recovered. The decrease in the recovery of NFG and entrained gangue per unit water was attributed to the coagulative nature of gangue in the presence of highly concentrated electrolytes and CMC. The fact that total NFG recovery did not change with water quality at a fixed CMC dosage but decreased with increasing ionic CMC dosage is indicative of the strong susceptibility of NFG depression to CMC dosage to an extent that at hyper dosages such as 500 g/t, all NFG is depressed completely. However, given the relationship between solids entrained and water recovery, the total recovery of entrained gangue increased with increasing ionic strength due to increased volumes of water which reported to the concentrate at increasing ionic strength. It was also shown that there was no change in sulphide recovery with increasing ionic strength. This was indicative of preferential adsorption of CMC onto gangue at the conditions tested. Higher solids recoveries or mass pulls were largely due to increased gangue recovery, mainly entrained gangue, which increased with increasing ionic strength. It was thus postulated that at increased ionic strengths, CMC coagulated gangue particles whilst indirectly destabilising the froth and retarding the action of electrolytes on froth stability through the removal of froth stabilising NFG. In order to examine coagulation at increasing ionic strengths, settling tests were performed on a Merensky ore and on pure talc. The results showed reduced settling time with increasing ionic strength and increasing CMC dosage suggesting that in a flotation cell, highly concentrated electrolytes would assist in depression by enhancing the coagulation of gangue and thereby decreasing their floatability. This can be attributed to increased concentrations of Ca2+ and CaOH+ which adsorb onto gangue; adsorption of which is the mechanism through which the chemisorption of CMC onto gangue occurs. In considering the effect of ionic strength and CMC dosage on froth stability, three phase froth column test results showed that the froth collapse time and froth height increased with increasing ionic strength due to an increase in the concentration of inorganic electrolytes which inhibit the coalescence of bubbles. In fact, Ca2+, Cl- , Mg2+, Na+, NO3 - and SO4 2- , which are present in the tested synthetic plant waters, are all reported in literature to have the ability to retard bubble coalescence, thus additive interactive effects in the tested systems should have been present. It was further shown that the addition of CMC resulted in a froth destabilization. The coagulation findings suggested that the presence of inorganic electrolytes enhanced the adsorption of CMC onto gangue due to changes on the mineral surface charge imposed by inorganic electrolytes. Microflotation results in the presence of CMC showed a decrease in the recovery of talc with increasing ionic strength whilst the presence of CMC did not affect the flotation behaviour of pyrrhotite. The adsorption results agreed with the microflotation results and the coagulation findings in that there was less residual CMC, meaning that more CMC was adsorbed onto the mineral surface with increasing ionic strength of plant water. In line with these findings, it was shown that the zeta potential of minerals, both talc and pyrrhotite, although investigated separately, increased (i.e. became less negative) with increasing ionic strengths. Thus, this work showed that increasing the ionic strength of plant water increased the concentration of inorganic electrolytes present in process water which acted on the mineral surface, passivating the mineral surface as seen through the less negative zeta potential in high ionic strengths. This would in turn create an environment conducive for an acid-base interaction between the hydroxyl species coated mineral particles (base) and highly negatively charged CMC ligand (acid), enhancing the preferential adsorption of CMC onto gangue as shown by the increase in the absorbed CMC concentration onto talc. The increased CMC adsorption would consequently assist in the formation of CMC-gangue mineral flocs with an induced coagulative and hydrophilic nature as shown by the shorter settling time in increasing ionic strength. Further investigations were carried out with single salts of cations and anions common in process water in order to identify whether there were any ions with the greatest froth stabilising action and gangue depression; Although Sulphide recoveries did not change with specific ions, the sulphide grades were affected by ion type owing to changes in gangue recoveries. Sulphide grades were higher with divalent ions compared to monovalent ions. It was also shown that salts containing NO3 - resulted in the lowest froth stability, as indicated by water recoveries and froth collapse time, compared to those which contained SO4 2- and Clin solution. Ca2+ and SO4 2- resulted in the highest froth stability compared to Na+. This can be attributed to a better inhibition of bubble coalescence in divalent ions compared to monovalent ions. The divalent Ca2+ and Mg2+ resulted in the lowest gangue recoveries compared to the monovalent Na+.NO3 - resulted in the least gangue recoveries compared to SO4 2- and Cl- . These findings suggested an enhanced hydrophilic nature onto gangue by divalent cations than monovalent cations with an even greater impact in NO3 - containing solutions- . Similarly, coagulation measurements showed an enhanced coagulation in NO3 - compared to SO4 2- with greater coagulation achieved in Ca2+ compared to Na+. An increase in the order of Ca(NO3)2>CaSO4>NaNO3>Na2SO4 in the zeta potential of talc and pyrrhotite was seen. This supported the enhanced coagulation and depression in Ca2+ and NO3 - containing systems. Thus, the findings of this work offer an opportunity to better understand water quality effects on flotation and their implications on froth stability and gangue management. Also, it has been shown that specific ion effects on froth stability and gangue management exist. Overall this study has shown that bench scale flotation techniques such as batch flotation, froth column flotation and microflotation can be used to understand the effect that water quality can have on a specific ore or mineral and that such techniques can be complemented with established surface chemistry laboratory techniques such as adsorption, coagulation and zeta potential to understand the interactions occurring in the air-water, air-solids and solids-water interfaces responsible for a particular flotation performance.. Through lower gangue recoveries, improved coagulation, increased adsorption and zeta potential, it can be concluded that the divalent Ca2+ is most likely to improve gangue depression and even more so in the presence of CMC compared to monovalent Na+. Moreover due to its causing a reduction of bubble coalescence, Ca2+ could result in improved froth stabilities and less entrainment. The monovalent Na+ showed higher gangue recoveries but lower water recoveries due to its weaker froth stabilising action compared to the divalent Ca2+. The higher gangue recoveries could be attributed to entrainment, meaning that given the lower froth stabilising action, Na+ richer solutions are most likely to lead to higher entrainment of gangue. Through higher water recoveries and higher froth collapse time results, it has been shown that SO4 2- ions result in better froth stabilities compared to Cl- and NO3 - ions, and would thus need to be monitored carefully for the desired froth stability. Thus, this work demonstrated the role of inorganic electrolytes on CMC efficacy and gangue depression using adsorption, coagulation and zeta potential results. These results correlated well with this study’s batch flotation and microflotation results. Also, these showed evidence to the suggestions and deductions drawn out of the bench scale flotation results on the effects and mechanisms through which inorganic electrolytes affect gangue and froth stability. This study also demonstrated that the divalent Ca2+ had the greatest froth stabilising effect and the greatest depressive effect on gangue compared to the monovalent Na+. Moreover, it provided evidence suggesting that solutions containing NO3 - were depressive on gangue and less froth stabilising compared to SO4 2- and Cl- . Findings of this work showed experimental evidence of the nature of CMC-electrolyte interaction in the pulp phase and its implications on the froth phase and gangue depression. It is believed that findings of this work offer an opportunity for flotation operations to tailor or control their water quality towards a desired flotation outcome. It may be possible that in order to combat changes in water quality, should closed water cycles be implemented, an operation could adjust their reagent suite to obtain a manageable grade and recovery and alleviate the high cost associated with cleaning of on-site water.
Open Access
Investigating the effect of water quality on the adsorption of a xanthate collector in the flotation of a sulphide ore
(2018) Manenzhe, Resoketswe; Corin, Kirsten; Wiese, Jennifer; Manono, Malibongwe
Environmental concerns necessitate the recycling of process water within mining operations. On average, recycled water contains more dissolved solids than fresh water. Since water is used as both a transportation and process medium, it is expected that changes in its quality will affect plant processes. Flotation is a process that is acutely sensitive to the immediate conditions of the system. Literature suggests that the efficiency of flotation separation is driven by the hydrophobicity that can be achieved by the mineral particles meant to be floated. The hydrophobicity is in turn driven by the adsorption of specialised reagents i.e. the collectors. Since collectors are added such that they adsorb at the liquidparticle interface, it stands to reason that changing the chemical composition of the aqueous phase will affect the collector adsorption, and hence the flotation response of target minerals. In this study, a sulphide copper ore from the Zambian Copperbelt was floated in synthetic plant waters of varying ionic strengths, and with different concentrations of the collector sodium isobutyl xanthate (SIBX). The synthetic plant waters were prepared by adding varying concentrations of inorganic salts to distilled water in order to achieve process water compositions that reflect water compositions typically found in mining plants. Additionally, a nickel-copper ore from Lapland Finland was floated in the synthetic plant waters as well actual plant waters. To account for the latter ore’s polymetallic nature, the collectors aerophine and sodium isopropyl xanthate (SIPX) were used sequentially. The objective of the study was therefore to investigate the effect of water quality on collector adsorption in the flotation of sulphide ores. The study showed that water quality has a quantifiable effect on SIBX and SIPX adsorption. The respective effects of water quality and collector adsorption on ore flotation could not be irrefutably decoupled. However, it could be concluded that of the tested waters, the copper thickener overflow was the least conducive to xanthate adsorption and valuable mineral recovery. On the other hand, collector adsorption was favoured by waters such as the raw and standard process. However, increased adsorption did not necessarily result in improved grades and recoveries. The study further showed that in the case that the dissolved ionic species are identical, increasing the ionic strength of water yields a linear decrease in xanthate adsorption, and that recycling SIPX retained in flotation waters resulted in reduced separation selectivity.
Open Access
Investigating the Influence of the Electrochemical Environment on the Flotation of a Mixed Sulphide Mineral System of Bornite and Chalcocite
(2022) Tafirenyika, Tanaka; Corin, Kirsten; O'connor, Cyril
There is a growing demand for copper driven by its applications in renewable energy and electric vehicles. Sulphide ores are an important source of copper. These ores contain, on average, 2% copper and require extensive processing to extract this as pure copper. Flotation is a critical front-end process used to remove gangue minerals and concentrate the copper minerals. However, flotation is an electrochemically intense process with multiple redox reactions taking place simultaneously. The interdependency of these processes makes it extremely difficult to isolate the effect of one parameter, and hence it is difficult to predict flotation behaviour. The electrochemical activity of sulphide minerals contributes to the overall activity in the flotation pulp. This ability to conduct electrons is called their rest potential, and different sulphide minerals have different natural rest potentials and therefore different extents of activity. The ability to conduct electrons gives rise to galvanic interactions between different minerals, minerals and media or a mineral-media-mineral complexes in solution. This dictates how collectors interact with the mineral surface, and ultimately the flotation response. Other variables in the electrochemical environment such as the dissolved oxygen, (DO), pH, redox potential (Eh) and water composition are also essential in controlling flotation outcomes. The aim of this investigation is to determine the influence of the electrochemical environment on the flotation of two copper sulphide minerals: bornite and chalcocite. The focus areas include; collector-mineral interaction, surface charge and flotation recovery of the bornite and chalcocite under varying pH conditions. For this investigation, the interaction of collector with bornite and chalcocite are considered in 2 water compositions: synthetic plant water (SPW1) and deionized water (DIW) at 5 different pH levels, increasing from 3 to 11. Starvation dosages of sodium isobutyl xanthate (SIBX) are used both in the batch flotation and collector adsorption tests conducted. Zeta potential tests are carried out to determine the surface charge of the minerals under the varying conditions. Further to the pure mineral studies, batch flotation tests are carried out using a synthetic ore under natural pH, Eh and DO conditions, with grinding done in a Magotteaux Mill® to monitor pulp chemical conditions during milling. From pure mineral flotation studies, it was observed that high mineral recovery is possible in both acidic and alkaline conditions via different hydrophobicity inducing mechanisms. In acidic conditions, low pH, two possible processes are occurring, the first being the inhibition of the formation of oxy-hydroxy species that adsorb onto the mineral surface and block sites for collector adsorption. The second is the decomposition of xanthate resulting in the formation of carbon disulphide that is known to be hydrophobic and is speculated to form a film around the mineral surface and render it hydrophobic enhancing flotation. In alkaline conditions, the well-established mechanisms of xanthate ion adsorption and dixanthogen formation take place on the mineral surface and enhances the flotation. Thus, the surface charge of the minerals as the conditions changed from acidic to alkaline resulted in a change in the surfaceactive species from the pure mineral to the oxide. However, for bornite, owing to the mineral structure containing iron, when oxidation occurs, iron hydroxide species form which precipitate at the mineral surface and inhibit collector adsorption, reducing the floatability of the mineral. The surface charge of the minerals' changes with changing pH due to a change in the surface-active species from the pure mineral to the oxide and hydroxide. Flotation of chalcocite and bornite in a mixed mineral system resulted in higher copper recovery compared to the weighted sum recoveries of the individual minerals. This suggests a possible synergistic effect when floating chalcocite and bornite together. Electrochemically active impurities present in the mineral samples made it difficult to decouple the exact nature of the interaction between the two minerals, but nonetheless provides insightful observations as industrial operations have to process equally complex mineral systems.
Open Access
An investigation into the effect of ionic strength of plant water on valuable mineral and gangue recovery of a platinum bearing ore from the Merensky reef
(2012) Manono, Malibongwe Shadrach; Corin, Kirsten; Wiese, Jenny
Higher solids and water recoveries were obtained at higher ionic strength. The increase in the ionic strength in the absence of any depressant caused an increase in Cu and Ni recovery.
Open Access
An investigation into the relationship between electrochemical properties and flotation of sulphide minerals
(2016) Chimonyo, Wonder; Corin, Kirsten; Wiese, Jenny
There is a growing importance in the mineral processing industry to find ways which are economic and effective in improving the recovery of minerals in the flotation process. The focus of this study was on the recovery by flotation of minerals found in the Merensky reef, which is one of the major reefs in the Bushveld complex. In that reef, base metal sulphide (BMS) minerals are commonly associated with PGMs and this has an effect on the way in which these minerals are concentrated by flotation (Vermaak et al. 2004; Wiese et al. 2005b; Miller et al. 2005; Schouwstra et al. 2000).A major problem in this process has been reported to be losses of valuable minerals (PGMs) associated with the loss of BMS (Wiese et al. 2005b), during flotation. The present investigation has focused on studying the relationship between the flotation of sulphide minerals using xanthates as collectors and the electrochemical properties of the flotation system. It is well known that electrochemical mechanisms in flotation systems have a major influence on flotation since the reactions occurring at the mineral/solution interface are of critical importance in the process (Woods, 1971).The aim of this study was to investigate the extent to which there was a relationship between the electrochemical reactions occurring in this ore which could indicate the effectiveness of the flotation process. The electrochemical reactions were studied by determining the redox potential changes occurring when various changes were made. These were the length of the alkyl chain length of the xanthate collector, changing the pH or using various chemical reagents to change the potential of the system. It was found from the rest potential measurements, that collectors of different chain length have different extents of interaction with mineral surface. A greater interaction, which is indicated by a greater change in the mixed potential after addition of the collector, is considered to be indicative of a greater adsorption of the collector at the mineral surface. It was hypothesized that this stronger adsorption by collectors of longer alkyl chain length would result in improved flotation performance. However, this was not observed to be the case and that was consistent with previous results on the relationship between the recovery of sulphide minerals in the Merensky ore and xanthates of different chain lengths. Thus it was shown that there was no correlation between the interactions between collectors of different alkyl chain lengths as determined through electrochemical studies and the flotation performance of valuable minerals under the tests conditions used.
Open Access
An investigation into the role of DTP as a co-collector in the flotation of a South African PGM ore
(2011) Bezuidenhout, Jacques; Corin, Kirsten; O'Connor, Cyril
The primary aim of this study was to investigate, using a combination of batch flotation and ToF-SIMS experimental techniques, whether a collector-collector synergistic interaction between SIBX and diethyl DTP will result in significantly enhanced copper, nickel, platinum and/or palladium recoveries and grades in a PGM containing ore from South Africa.
Open Access
Investigation of the effect of the reagent suite in froth flotation of a Merensky ore
(2015) Moimane, Tiisetso Makheane; Corin, Kirsten; Wiese, Jenny
The mining industry is faced with a challenge to develop efficient and economically feasible processing routes owing to the depletion of high-grade ores, and the ever increasing demand for precious metals for a wide range of applications. The valuable minerals, PGMs (Platinum Group Minerals) and BMS (Base Metal Sulphides) in the ores are extracted through the aid of chemical reagents (activators, collectors, depressants, frothers, modifiers) which are added to the flotation circuits to facilitate the separation between these minerals and the undesired gangue minerals present in the ore. The process is made complex by many surface reactions taking place, the existence of secondary and interactive effects among the flotation reagents, as well as the surface liberation of the minerals. Owing to the stringent regulations around water usage, concentrator plants are left with no option but to recycle water within their operations. This practice leads to accumulation of pollutants, such as organics, flotation reagents residues, dissolved ions, etc., which will likely have an influence on the chemical environment of the process, and subsequently will bear an impact on the overall metallurgical performance of the concentrator. This makes the process even more intricate, making it difficult to account for the behaviour of the chemical reagents, as well as making it virtually impossible to precisely assess their individual contribution to the overall flotation performance. Hence it is of crucial importance to adopt a holistic approach when investigating the effects of the chemical parameters in a flotation process. This is a flotation chemistry study that adopted a two-level-four-factor (24) factorial experimental design to evaluate the simultaneous effects of the chemical parameters, with particular reference to collector, depressant, frother, and ionic strength of the synthetic plant water, as well as determining the possible interactive effects between the chosen parameters. This investigation was made possible by conducting batch flotation tests on a PGM-bearing ore from the Merensky reef of the Bushveld Igneous Complex. Sodium isobutyl xanthate (SIBX), a polysaccharide, namely guar gum, and a polyglycol ether, namely Dowfroth 250 were used as the collector, depressant and frother, respectively. These are the typical chemical reagents the dosages of which tailored in the PGM industry for the processing of the ores. The metallurgical performance indicators used were solids, water, copper and nickel recoveries as well as copper and nickel grades.
Open Access
Simulating the Effect of Water Recirculation on Flotation through Ion-Spiking: Effect of Ca2+ and Mg2+
(2020-11-19) Dzingai, Mathew; Manono, Malibongwe; Corin, Kirsten
Froth flotation is a multifaceted complex process which is water intensive. The use of recycled water as an alternative source of water to meet water demands of the process may introduce deleterious inorganic ions that affect the mineral surface, pulp chemistry, and reagent action, hence the need to establish whether threshold ion concentrations exist beyond which flotation performance will be adversely affected. This is of paramount importance in informing appropriate recycle streams and allowing simple, cost-effective water treatment methods to be applied. Here we report that increasing ionic strengths of synthetic plant water (SPW); 3, 5, and 10 SPW respectively, resulted in an increase in water recovery in the order 0.073 mol·dm⁻³ (3 SPW) < 0.121 mol·dm⁻³ (5 SPW) < 0.242 mol·dm⁻³ (10 SPW), indicating an increase in froth stability as higher water recoveries are linked to increased froth stabilities. This behavior is linked to the action of inorganic electrolytes on bubble coalescence which is reported in literature. There was, however, no significant effect on the valuable mineral recovery. Spiking 3 SPW to 400 mg/L Ca²⁺ resulted in higher copper and nickel grades compared to 3 SPW, 5 SPW, and 10 SPW and was deemed to be the Ca²⁺ ion threshold concentration for this study since 3 SPW spiked with further Ca²⁺ to a concentration of 800 mg/L resulted in a decrease in the concentrate grade. The spiking of 3 SPW with Mg²⁺ resulted in higher copper and nickel grades compared to all other synthetic plant water conditions tested in this study.
Open Access
The Solution Interaction of Tetrathionate Ions and Sodium Isobutyl Xanthate and Its Effect on the Flotation of Galena and Chalcopyrite
(2021-02-15) Mhonde, Ngoni; Pitkänen, Leena; Corin, Kirsten; Schreithofer, Nóra
Tetrathionates have been found in significantly high concentrations in recycled process waters from massive sulphide ore processing plants. These polythionates react with xanthate added to flotation pulps thus reducing xanthate dosages in solution potentially affecting flotation performance. The current study focused on the effect of the tetrathionate-xanthate reaction on sulphide mineral recoveries. Ore dissolution studies confirmed the generation of tetrathionates by copper-lead-zinc ores. In 20 min, the tetrathionates consumed more than half of the xanthate in solution at pH 7. Rest potential measurements and Fourier transform infrared spectroscopy (FTIR) showed that the degree of collector-mineral interactions of xanthate and both galena and chalcopyrite was greatly reduced in the presence of a 2000 mg/L tetrathionate solution. Microflotation tests showed that chalcopyrite recovery was less sensitive to tetrathionates as indicated by small changes in mineral recoveries. Galena was sensitive to the action of tetrathionates on the mineral surface as the galena recovery significantly declined when floated with xanthate as a collector in both a 500 mg/L tetrathionate solution and a 2000 mg/L tetrathionate solution. These fundamental results lay a sound base on which more discussion into the significance and the effect of tetrathionates on flotation performance of sulphide ores can be developed.
Open Access
A study of the flotation characteristics of a complex copper ore
(2015) Kalichini, Monica Shamvuse; Corin, Kirsten; O'Connor, Cyril
Kansanshi copper mine is situated in the north western province of Zambia. Weathering has given rise to a vertically zoned profile comprising leached, refractory, oxide, mixed and hypogene sulphide mineralisation. As a result of the mineral variations, the processing plant treats three distinct ore types; oxide, sulphide and mixed. The objective of this study was to investigate the floatability of a complex Kansanshi mixed copper ore comprising sulphide and oxide minerals with a view to achieving an optimal flotation performance in the treatment of the Kansanshi ore body. This required an in-depth analysis of the mineralogy of the feed as well as tailings samples after different flotation processes involving a range of reagent types and dosage procedures. The ore samples studied represented a high quality (HQ) ore dominated by sulphide minerals and low quality (LQ) ore dominated by oxide minerals. The quality of an ore at Kansanshi is defined by the acid soluble copper (ASCu) content of the ore, which is used as a proxy for oxide mineral content. An important finding in this study was that sulphide minerals are also prone to digestion during this analysis. Chalcopyrite was the major copper mineral in the HQ ore, constituting 3.9 %, but only 1.0 % of the LQ ore. LQ ore was dominated by chrysocolla, which constituted 3.8 % of the ore. The treatment of HQ ore with 30 g/t SIBX has shown that up to 90 % of the copper can be recovered from HQ ore. On the other hand, while 30 g/t SIBX was sufficient for chalcopyrite recovery in LQ ore, the tailings mineralogy after flotation with SIBX indicated that 78.8 % of the unrecovered copper in LQ ore was present as chrysocolla, 1.3 % as malachite and 5.8 % as chalcopyrite and therefore LQ ore required alternative flotation methods for the recovery of the oxide minerals. Comparison of slug sulphidisation and controlled potential sulphidisation (CPS) of LQ ore have shown that CPS performs better than slug sulphidisation only when the correct potential range and SIBX dosage after sulphdisation are used. Tailings mineralogy of LQ ore after sulphidisation showed a copper deportment of 0.1 % cuprite, 0.6 % malachite, 0.8 % chalcopyrite and 84.8 % chrysocolla, suggesting that all oxide copper minerals present in the LQ ore, except chrysocolla, are amenable to flotation using SIBX after sulphidisation. This observation was further verified through sulphidisation in a microflotation cell, which showed malachite recovery of 18.2 % compared to only 0.5 % of the chrysocolla. A techno-economic analysis comparing slug sulphidisation and CPS has indicated that CPS using a potential range of -300 to -400 mV performs better than slug sulphidisation from an economic stand point. At this potential, a NaHS:SIBX ratio of 7:1 was required, further highlighting the importance of using the correct collector dosage after sulphidisation.

Browsing by Author "Corin, Kirsten"

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