Search Results (19)

The reprocessing of metallurgical wastes to recover much-needed metals such as copper not only ensures an adequate supply of metals but also contributes to the cleaning of the environment. A copper smelter in Namibia accumulated significant amounts of spent refractory bricks that are enriched with metal values including copper. This supposedly waste material can potentially serve as a supplement to the ore concentrate, as a smelter feedstock for this toll smelter. Representative samples of crushed bricks, designated as Sample 1 and Sample 2, were used for mineralogical characterisation and flotation test work. The assays for Sample 1 and Sample 2 were 14% Cu and 18% Cu, respectively. Microscopy results identified various copper phases including metallic Cu, bornite, malachite and chalcopyrite. Batch flotation tests were conducted to investigate the effect of grind size (P₈₀ of 53, 75 and 106 μm), pulp pH (natural pulp pH, 10, 10.5 and 11) and collector (potassium amyl xanthate, PAX) dosage (70, 100 and 130 g/t) on the recovery of copper, concentrate grade and weight recovery. In some tests, a co-collector (dithiophosphate, DTP) and sulphidiser (Na₂S) were also added in the quest to maximise the recovery of copper. Based on the test conditions investigated in this study, the grind size is the key variable affecting the recovery of copper. The best copper recovery of 86% (with a weight recovery in the range of 42 to 45% (w/w) and concentrate grade of 37% Cu) was achieved for the finest grind size of 53 μm. The reagent suite that yielded the best recovery was 70 g/t PAX with no addition of the sulphidiser while the pH was 10. There is scope for developing the process routes to recover other valuable metals such as iron, lead and zinc that are also in the spent bricks, as well as potential reuse of the spent bricks (after recovering valuable metals) to make new refractory bricks. Full article

This study demonstrated the successful recovery of zinc, lead, and copper collective concentrates from historical metal-bearing mine tailings (sulfide–polymetallic ore with a composition of 7.38% Zn, 1.45% Pb, and 0.49% Cu) using froth flotation techniques, which were originally developed during uranium ore mining. Froth flotation techniques were used to justify suitability for recovering metals. The effects of a dosage of the foaming agent Polyethylene glycol (PEG 600) at 50 and 100 g t⁻¹, collector types Aerophine 3418A (AERO), Danafloat 067 (DF), and potassium ethyl xanthate (KEX) at 50 and 80 g t⁻¹, and a suspension density of 300 and 500 g L⁻¹ on froth flotation collective concentrates were investigated. The final collective concentrate achieved recoveries exceeding 91% for lead (Pb), 88% for copper (Cu), and 87% for zinc (Zn). The obtained concentrates were analyzed using Atomic Absorption Spectroscopy (AAS) and X-ray Fluorescence Spectrometry (XRF), while selected samples were further examined via Scanning Electron Microscopy (SEM) with Energy Dispersive X-ray Spectroscopy (EDS). The resulting sulfide concentrates can subsequently be treated using suitable hydrometallurgical techniques. The application of these concentrates in metal production would help reduce the environmental burden of mining activities. Full article

Millions of tons of zinc plant leach residues (ZPLR) have been stockpiled in Iranian hydrometallurgical zinc plants during the last few decades. Due to the low grades of zinc, lead, and silver in these residues, these residues have been abandoned without treatment. The authors of this paper studied zinc plant leach residues (ZPLR) to propose a flotation process for separating and producing lead and silver concentrate. A response surface methodology (RSM) was employed to obtain six models for optimizing the best conditions for lead recovery, lead grade, zinc recovery, zinc grade, silver recovery, and silver grade. In these models, the effect of the different main variables, including density, flotation time, pH, sodium sulfide dosage, and potassium amyl xanthate dosage, was investigated to optimize grades and recoveries. The studied ZPLRs were categorized into two types based on the disposal time, including new and old residues. The chemical analysis showed that the grades of lead, zinc, and silver in the new residues are higher than in the old residues. In a previous mineralogical study, it was found that silver forms in lead and zinc minerals as a solid solution within their structures. The resulting 3D graphs showed that the interacting variables have significant effects on responses. The ANOVA analysis exhibited the order of model significance to be lead grade (F-value of 36.46) > silver grade (19.76) > lead recovery (7.88) > zinc grade (5.63) > silver recovery (5.58) > zinc recovery (4.83). Based on these models, under the conditions of 1126.26 g/cm³ density, 20.83 min retention time, 9.9 pH, 6 kg/t sodium sulfide, and 749.66 g/t potassium amyl xanthate dosage for a new residue type, the recoveries of lead, zinc, and silver were determined to be 51.10%, 11.13%, and 72.85%, with grades of 38.87% Pb, 8.46% Zn, and 1209.11 g/t Ag, respectively. According to the feasibility study results, the presented work is reasonable in terms of technical, economic, and investment potential. Full article

The sulfidization-amine flotation method is commonly used for the beneficiation of zinc oxide ores. Lanping zinc oxide ores contains 8.40% zinc, with the main mineral being smithsonite; additionally, they have a high mud content. Conventional sulfidization–ammonium flotation presents a low flotation index and unsatisfactory flotation froth. A new open-circuit technology is employed to treat Lanping zinc oxide ores, where Na₂S, KG-248, and dodecyl amine + sodium isoamyl xanthate + ammonium dibutyl dithiophosphate are used as the regulator, depressant, and ternary collector, respectively. Consequently, the flotation indices for the zinc grade and recovery are 28.71% and 86.24%, respectively, and the flotation froth becomes more stable. Subsequently, the flotation behavior and adsorption mechanism of smithsonite with a ternary collector are investigated. The flotation recovery of smithsonite increases to 94.40% after treatment with the ternary collector. Surface-analysis results indicate that the ternary collector can synergistically adsorb onto the sulfidized smithsonite surface to enhance its hydrophobicity, thus increasing the floatability of smithsonite. Meanwhile, the total consumption of the collector in the ternary-collector system is lower than that in the binary- or unitary-collector system. Full article

The use of toxic chemicals as flotation reagents raises concerns about the sustainability of the process and its environmental impact. Xanthates are a family of efficient collectors that find application in the selective recovery of lead, gold and zinc ores, but they are toxic, with the majority of them being imported from eastern countries. Recently, organosolv lignin (OLN) micro- and nanoparticles have been produced and tested as collectors in the flotation of Cu minerals with positive results. The material is attractive because it is natural, biodegradable, and possesses a low carbon footprint compared to the conventional reagents. A mixed sulphide ore deposit in Halkidiki, Greece, is operated by Hellas Gold S.A., a subsidiary of Eldorado Gold. The company produces galena (PbS), sphalerite (ZnS) and Au bearing pyrite/arsenopyrite (FeS₂/FeAsS) concentrates by applying consecutive froth flotation stages. Our previous laboratory studies have shown that the partial substitution of sodium isopropyl xanthate (SIPX) with OLN is possible, without worsening the flotation performance or downgrading the concentrates’ quality. This study presents the findings of 3-stage pilot-scale flotation tests using OLN as a partial substituent of the SIPX collector on sphalerite and pyrite/arsenopyrite circuits. In the sphalerite recovery circuit, the partial replacement of SIPX with OLN (25 and 50%) resulted in an increase in Zn grade and a similar recovery compared to the standard case, while better selectivity was achieved since the Au recovery in the Zn concentrate was lower. Similarly, with the pyrite/arsenopyrite flotation circuit, the replacement of SIPX with OLN resulted in an increase in gold recovery with a parallel reduction in Pb recovery. It appears that OLN can efficiently replace part of the SIPX collector in Zn and Au flotation, producing concentrates of similar to better purity, in terms of grade and recovery of valuable metals, because of the improved selectivity of the mixed collector. The introduction of OLN in the collector mixture and the parallel reduction of SIPX drastically reduce the environmental footprint of the process. Full article

Sphalerite is a common sulfide mineral, and xanthate is usually used as the collector of sphalerite flotation. In order to realize the simultaneous flotation of sphalerite- and zinc-containing oxide ores, in this paper, dodecylamine was used as the collector and sodium sulfide as the regulator to explore the flotation behavior of sphalerite in a sulfuration–amine system through single mineral flotation tests, and several analytical methods, including contact angle measurement, adsorption capacity test, FT-IR, and XPS, were employed to discuss the interaction mechanism of dodecylamine on the surface of sphalerite. The results confirm that dodecylamine had a strong collecting effect on sphalerite throughout almost the entire pH range, and the flotation recovery increased with the increase in pH, reaching 86.81% at pH 12. It was found that the effect of sodium sulfide on the floatability of sphalerite was relatively weak, and a weak activation effect, of about 2% improved recovery, was achieved. The results of the mechanism analysis further show that dodecylamine underwent both physical adsorption and chemical adsorption on the surface of sphalerite, and the physical adsorption was enhanced with the increase in pH. The above findings provide evidence for the simultaneous flotation of zinc-containing sulfide-oxide mixed minerals. Full article

Mining and mineral processing industry adversely affects ecosystems and communities in nearby areas, including high freshwater consumption and scarcity. That is why the emerging global trend is to use sea water in flotation to recover valuable minerals from finely disseminated base metals ores. Recent studies investigate sea water flotation of copper, molybdenum, nickel sulphides and pyrite, while flotation of sphalerite, the main valuable mineral for zinc production, remains uncovered. This paper examines the feasibility of sphalerite flotation by conventional collectors in artificial sea water using a bubble-particles technique and frothless flotation tests. Potassium isopropyl xanthate (PIPX) and sodium isopropyl dithiophosphate (SIDTP) were used as collectors, and copper sulphate was introduced as the activator, while zinc sulphate and sodium sulphide were used as depressants. We examined the most common size fractions of sphalerite: medium (−74 + 44 μm) and fines (−44 μm). The findings showed the feasibility of sphalerite flotation in artificial sea water. We also established correlations between the rate of bubble-particle attachment and the sphalerite flotation recovery resulting in the growth of flotation recovery with the increase of the bubble-particle attachment rate. The results can be used as guidelines in choosing flotation reagents for sphalerite flotation in sea water. Another practical application of the results is the potential for sustainable development of the industrial sector, ecosystems and societies due to the replacement of fresh water by sea water, although further technological and environmental studies are required. Full article

The lead–zinc ore is a typical low-grade complex polymetallic sulphide ore. Its typical mineralogy is characterized by a high content of cryptocrystalline graphite, low content of lead, zinc, and copper, and fine grain size. However, the aggregation size of sulfide mineral aggregations is coarse, and the symbiotic relationship is relatively simple. Conventional process treatment requires the fine grinding of ore into mineral monomer dissociations followed by preferential flotation, but this method has a poor sorting index and high production cost. In this paper, the grain size of the sulfide mineral aggregates determined by mineralogical studies is used to determine the fineness of the grinding, so that the liberation degree of the sulfide ore in the coarse grinding product reaches 70%, and each flotation process is used to collect rough concentrate. In the first step of flotation, the carbon in lead–zinc sulfide ore was removed by adding #2 oil to the flotation tank at a dosage of 40 g/t. In the second flotation process, the pH was adjusted to 7.2, the dosage of isopropyl xanthate was 20 g/t, the dosage of #2 oil was 30 g/t; the flotation reagent in the third step was isopropyl xanthate, and the dosage was 7.5 g/t, with a #2 oil of 15 g/t. Each flotation process concentrate is collected. Finally, the grades of lead and zinc in the final concentrate were 1.6% and 5.71%, respectively. In addition, the recoveries of lead and zinc were 91.78% and 92.07%, respectively. The yield of tailings was 50.6%. By the flotation of sulfide aggregates, a large number of gangues are avoided to participate in fine grinding, which helps to reduce the energy consumption of the mill. Full article

Reuse of process water in the flotation of sulfide ores has crucial importance for environmental sustainability and improved process economy. However, the chemistry of process water may be a critical subject for the flotation efficiency as the dissolved ion concentration increases with water reuse. In this study, the effects of water reuse on the flotation efficiency of a Cu-Zn sulfide ore were investigated. The flotation flowsheet consists of a pre-flotation section to remove the naturally floatable talc particles, and sequential copper and zinc flotation sections. Calcium, sulfate, thiosulfate, and xanthate were found as major contaminating ionic species in the process water discharged from flotation circuits. The flotation tests with recirculated water from the zinc rougher tailing revealed that the presence of residual xanthate caused unintentional activation of copper minerals in the pre-float section. Copper recovery increased in the pre-flotation section and resulted in the loss of copper to the pre-float concentrate, which is considered as a tailing stream in the current flowsheet. Various types of activated carbon samples were tested to remove the residual xanthate from the tailing water. The carbon samples could be regenerated by heat treatment and reused for water treatment. Performance of the activated carbon samples was directly related to the pore size and surface area. Carbon-treated tailing water could be re-used in flotation without affecting the flotation performance. Full article

The use of untreated recycled water has negative effects in the flotation of zinc sulfide ores due to the presence of dissolved species, such as magnesium and calcium. Although it has been found that magnesium is a more potent depressant than calcium, it has not been investigated in this role or for the effect of adding sodium carbonate. The results of an investigation to evaluate the effect of magnesium on the hydrophobicity of Cu-activated sphalerite conditioned with Sodium Isopropyl Xanthate (SIPX) are presented. Zeta potential of natural and Cu-activated sphalerite as a function of the conditioning pH and Cu(II) concentration, respectively, was first evaluated. Later, the effect of pH and presence of magnesium on the contact angle of Cu-activated sphalerite conditioned with SIPX was studied; it was also evaluated the effect of sodium carbonate to counteract the effect of magnesium. Cu-activation enhances the zeta potential of sphalerite up to a concentration of 5 mg/L. Contact angle tests, thermodynamic simulation, and surface analysis showed that magnesium hydroxide precipitates on the sphalerite surface at pH 9.6, decreasing its hydrophobicity. Addition of sodium carbonate as alkalinizing agent precipitates the magnesium in the form of a species that remained dispersed in the bulk solution, favoring the contact angle of Cu-activated sphalerite and, consequently, its hydrophobicity. It is concluded that the use of sodium carbonate as alkalinizing agent favors the precipitation of magnesium as hydromagnesite (Mg₅(OH)₂(CO₃)₄∙4H₂O) instead of hydroxide allowing the recovery of sphalerite. Full article

The specific objectives of the presented study were related to the optimization of the production process of N-alkyl-, N,N-dialkyl-, and N-cycloalkyl-O-isobutyl thiocarbamate; trial industrial production of N-ethyl-O-isobutyl thiocarbamate; and the evaluation of flotation efficiency of N-ethyl-O-isobutyl thiocarbamate using a real ore sample. The optimization of thiocarbamate syntheses were performed by varying the molar ratio of isobutyl alcohol, carbon disulfide, potassium hydroxide, reaction time, and reaction temperature. In the first step, one-pot reaction took place to produce alkyl xanthate and was followed with chlorination to give alkyl chloroformate (O-alkyl carbonochloridothioate); finally, thiocarbamates were obtained by the reaction with corresponding amines. N-alkyl-O-ethyl thiocarbamate was synthesized as a comparative flotation agent. The structure of the synthesized compounds was confirmed by IR, ¹H and ¹³C NMR, and MS instrumental methods, and the purity was determined by gas chromatographic method and elemental analysis. The optimized methods gave high-purity products in a significant yield that was also confirmed by semi-industrial production of N-ethyl-O-isobutyl thiocarbamate. The optimized thiocarbamate synthesis, without isolation of intermediates, is of great importance from the aspect of green technologies. Flotation efficiency test results, using real copper and zinc ores, showed the highest activity of N-ethyl-O-isobutyl thiocarbamate. The optimal one-pot thiocarbamate synthesis provides a simple procedure with a high conversion degree, and, thus, offers valuable technology applicable at the industrial scale. Full article

Heavy metals pollution receives worldwide attention due to great toxicity, significant bio-accumulation and non-biodegradability. Adsorption is a promising technique for removing heavy metals from wastewater. Adsorption of zinc (Zn(II)) from aqueous solution was investigated by functionalized lignocellulose derived from fallen leaves. Alkalized lignocellulose (AC), xanthated lignocellulose (XC) and carboxylated lignocellulose (CC) were characterized by Fourier transform infrared spectrum (FTIR), X-ray photoelectron spectroscopy (XPS) and scanning electron microscopy (SEM). The effect of sorbent dosage, solution pH, sorption time and initial Zn(II) concentration on Zn(II) sorption was investigated by single-factor experiment. Sorption kinetics, isotherms and thermodynamics were examined to reveal sorption mechanism. The sorption capacity and removal rate remarkably depend on experimental variables. Zn(II) sorption onto AC, XC and CC is well described by the pseudo second order kinetics and Langmuir isotherm. The sorption process is fast, reaching sorption equilibrium at 30 min. The maximum sorption capacity of Zn(II) onto CC is 46.49 mg/g, higher than that onto AC, XC and other reported sorbents. Thermodynamic parameters indicate that Zn(II) sorption is a spontaneous process. Sorption mechanism is majorly attributed to surface complexation. This work shows the feasibility of removing toxic Zn(II) from aqueous solution by locally available biomass, providing a sustainable approach for wastewater treatment. Full article

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. Full article

The purpose of this study is to propose the flotation procedure of seafloor massive sulfide (SMS) ores to separate chalcopyrite and galena as froth and sphalerite, pyrite, and other gangue minerals as tailings, which is currently facing difficulties due to the presence of water-soluble compounds. The obtained SMS ore sample contains CuFeS₂, ZnS, FeS₂, SiO₂, and BaSO₄ in addition to PbS and PbSO₄ as Pb minerals. Soluble compounds releasing Pb, Zn²⁺, Pb²⁺, and Fe^2+/3+ are also contained. When anglesite co-exists, lead activation of sphalerite occurred, and thus sphalerite was recovered together with chalcopyrite as froth. To remove soluble compounds (e.g., anglesite) that have detrimental effects on the separation efficiency of chalcopyrite and sphalerite, surface cleaning pretreatment using ethylene diamine tetra acetic acid (EDTA) was applied before flotation. Although most of anglesite were removed and the recovery of chalcopyrite was improved from 19% to 81% at 20 g/t potassium amyl xanthate (KAX) after EDTA washing, the floatability of sphalerite was not suppressed. When zinc sulfate was used as a depressant for sphalerite after EDTA washing, the separation efficiency of chalcopyrite and sphalerite was improved due to deactivation of lead-activated sphalerite by zinc sulfate. The proposed flotation procedure of SMS ores—a combination of surface cleaning with EDTA to remove anglesite and the depression of lead-activated sphalerite by using zinc sulfate—could achieve the highest separation efficiency of chalcopyrite and sphalerite; that is, at 200 g/t KAX, the recoveries of chalcopyrite and sphalerite were 86% and 17%, respectively. Full article

The effects of ZnSO₄ on arsenopyrite depression were studied with sodium carbonate and sodium isobutyl xanthate (SIBX) as the pH regulator and collector, respectively. In both micro and real ore flotation tests, ZnSO₄ showed better depression on arsenopyrite (pH 7.5–9.0 adjusted by Na₂CO₃) compared with sodium humate. The depression mechanism of ZnSO₄ on arsenopyrite flotation was studied by electrokinetic potential, adsorbed amount measurements, scanning electron microscope (SEM) observation and energy dispersive spectra (EDS) detection. The electrokinetic potential measurement results show a potential increase forpleas the arsenopyrite treated with ZnSO₄ in the pH range 7.5–9.0, which could be attributed to the formation of the precipitated zinc carbonate (ZnCO_3(S)). For arsenopyrite treated with both ZnSO₄ and SIBX, the electric surface potentials also display an increase, to approximate the values with solely ZnSO₄ treated, at pH 7.5–9.0, indicating the inhibition of ZnCO_3(S) upon the SIBX adsorption onto arsenopyrite. Adsorption results demonstrated that SIBX adsorption onto arsenopyrite indeed was inhibited at the pH 7.5-9.0 through the sharp decrease in SIBX adsorbed amount with ZnSO₄ as the depressant at this pH range. SEM observation and EDS detection results verify the formation of colloidal ZnCO₃ on the arsenopyrite, with ZnSO₄ as the depressant in combination with Na₂CO₃. Full article