Flotation Of Sulfide Minerals Research Articles

Insight into the effect of ultrasonic treatment on surface properties and flotation behavior of chalcopyrite after galvanic corrosion of grinding media and chalcopyrite

Ultrasonic treatment is an effective means to promote chalcopyrite flotation. Chalcopyrite needs to be grinded before flotation production. During the grinding process, galvanic corrosion occurs between chalcopyrite and iron grinding media. Galvanic corrosion significantly affects the efficacy of ultrasonic treatment on chalcopyrite flotation. Nonetheless, there is still a lack of research on ultrasonic treatment of chalcopyrite after galvanic corrosion. In this study, micro-flotation tests, SEM-EDS, TOF-SIMS, XPS, ICP-OES and electrochemical tests were used to study the effects of ultrasonic treatment on the surface properties and flotation behavior of chalcopyrite before and after galvanic corrosion under different ultrasonic conditions and galvanic corrosion conditions. The results show that in the absence of galvanic corrosion, short time and low power ultrasonic treatment can improve chalcopyrite floatability. However, when the chalcopyrite after strong galvanic corrosion is subjected to ultrasonic treatment, the increase of galvanic corrosion temperature, galvanic corrosion time, ultrasound power and ultrasound time result in a marked reduction of the flotation recovery of chalcopyrite. The synergistic effect of ultrasonic and galvanic corrosion significantly inhibits the flotation recovery of chalcopyrite, which may provide a new potential way for the inhibition of sulfide ore in flotation production. The research results are helpful to expand the use range of ultrasonic processing for sulfide mineral pretreatment and flotation, and provide a theoretical reference for the application of ultrasonic treatment in sulfide minerals with galvanic corrosion.

The influence of NaCl on xanthate adsorption on chalcopyrite surface and chalcopyrite flotation

Due to freshwater scarcity, mineral flotation plants have been using alternative water resources including seawater together with water recycling resulting in the accumulation of salts in process water. However, how salts affect the adsorption of xanthates on the surfaces of sulphide minerals and subsequent sulphide mineral flotation is unclear. To address the knowledge gap, this study investigated how NaCl affected the adsorption of potassium amyl xanthate (PAX) on chalcopyrite surface and chalcopyrite flotation. Cyclic voltammetry (CV) tests identified that NaCl increased the oxidation of PAX and its adsorption on chalcopyrite surface due to elevated ionic strength and electrical conductivity of water, facilitating charge transfer among PAX, chalcopyrite surface and oxygen. NaCl also increased chalcopyrite surface oxidation to produce more CuS, facilitating the formation of copper-xanthate complexes. Flotation tests identified that the increased adsorption of PAX by NaCl was only translated into improved true flotation of chalcopyrite at a low PAX concentration such as 50 ppm in NaCl solutions with low and medium concentrations (0.1 and 0.7 M, respectively). At a high PAX concentration such as 150 ppm, the coexistence of high concentrations of PAX and NaCl led to strong froth destabilisation, resulting in a dramatic reduction in water recovery and true flotation recovery of chalcopyrite. The findings from this study will be able to guide the flotation plants to maximise mineral flotation by choosing collector dosage based on the ionic strength of process water.

Variability of lead-zinc sulfide ore slurry rheological properties and its influence on flotation conditions

We investigated how the rheological characteristics of a flotation slurry change in response to variations in mineral species, slurry concentration, and particle size; slurry pH; and trap and rheological control reagent concentrations using slurries containing galena, sphalerite, quartz, and kaolinite. The results indicate that reducing particle size and increasing slurry concentration leads to varying degrees of increase in apparent viscosity and yield stress. At the same particle size, the slurry exhibits the following order of apparent viscosity and yield stress: kaolinite > galena > sphalerite > quartz. In addition, as the slurry’s apparent viscosity and yield stress increase, the rheology decreases, creating progressively unfavorable conditions for the flotation of lead, zinc, and other target minerals. Furthermore, changes in pH have no significant effect on the slurry’s rheology when the slurry is comprised of vein minerals. Moreover, galena and sphalerite depict particle agglomeration in the slurry. Ultimately, the addition of sodium silicate as a rheological control reagent substantially enhances the slurry’s rheological properties. This results in a system where problematic minerals like kaolinite are more effectively dispersed, thereby promoting efficient lead-zinc mineral flotation. Regarding the flotation of lead sulfide and zinc minerals, the addition of kaolinite raises the apparent viscosity of the mixed slurry, hindering the flotation of the target minerals. Conversely, quartz lowers the apparent viscosity aiding the flotation separation process. Understanding the relationship between flotation conditions and the pulp’s rheological properties can provide valuable guidance for subsequent flotation tests.

Enhancing the Flotation Separation of Galena and Pyrite in the Presence of Seawater.

The effect of seawater on the flotation of sulfide minerals is a hot research topic at the present. Using seawater instead of fresh water to separate lead-sulfur can save fresh water resources and reduce lime consumption, which has broad prospects in terms of economic and environmental protection. Many studies have shown that divalent calcium and magnesium ions in seawater are hydrolyzed to form hydroxyl complexes under alkaline pH values, which affects the flotation separation of sulfide ore. In this paper, the flotation separation of galena and pyrite in seawater was tested, and the mechanism was analyzed. Compared with that in deionized water, the lead grade of concentrate in seawater flotation increased from 57.40 to 60.20%. Meanwhile, the recovery of galena declined from 68.66 to 61.17%, and the SI value increased from 2.020 to 2.072. With the addition of SHMP, SS, and EDTA, the SI value increased from 2.072 to 2.609, 2.525, and 2.287, respectively. Noteworthily, when EDTA was excessive, the grade of Pb was 69.60%, and the SI value was 3.050. Wettability analysis, zeta potential analysis, X-ray photoelectron spectroscopy, Fourier transform infrared spectroscopy, density functional theory calculation, and E-DLVO calculation verified the microflotation results and revealed the mechanism of action.

Linking ores flotation results to the Kinsenda copper deposit mineralogical features: enhancing the recovery of copper retained in flotation tailings

ABSTRACT Since the Kinsenda concentrator commissioning in 2015, the selective flotation of mixed copper ores (67% as sulphides and 39% as oxides) from the Kinsenda deposit (DR Congo results in a recovery yield of 57% of copper as oxides versus 98% as sulphides.. Talings from the flotation of sulphide minerals contain 1.14% as total Cu including 0.38% as oxides. This research adopted a geometallurgical approach for enhancing the recovery of not floated copper. Consequently, it thoroughly studied the Kinsenda copper deposit mineralogy to explain causes of poor results of the industrial flotation of copper and conducted chemical and granulochemical analyses of tailings to describe and comprehend the copper behaviour during flotation. The effects from particles size in tailings, and the reagents’ suite components and their dosages on the copper flotation efficiency were investigated. Flotation results were significantly improved when tailings contain 70% particles with size smaller than 106 μm were floated using a reagents suite including 3600 g per t NaHS (sulphidiser), 360 g per t SNPX (collector) and 360 g per t AM810 (promoter). Indeed, the concentrate's grade in total copper reached 23.75%, with a recovery yield of 91.07% and 15.41% in oxidised copper, with a recovery yield of 91.26%.

Технология переработки сухих отходов сорбции и растворов хлоридов сурьмы

Reducing the loss of valuable components during the processing of complex gold-antimony ores, increasing the extraction of antimony during the flotation of sulfide minerals, is an urgent scientific problem. The aim of the study is to maximize the extraction of gold and antimony from dry sorption waste after gold cyanidation, and to improve the reagent regime of the antimony mineral flotation process. Research objectives are as follows: evaluation of the technology efficiency for processing dry sorption waste; extraction of gold from the cake of acidic leaching of antimony; production of various antimony-containing products from solutions of antimony chlorides; study of the possibility of replacing lead with a mixture of zinc and copper cations to hydrophobize the surface of antimony sulfide minerals during flotation. The object of the research is man-made and natural mineral raw materials containing gold and antimony. The following research methodology and methods are used: information analysis, evaluation of existing scientific developments, methods of theoretical and experimental research. A new technological solution has been proposed: acidic (a mixture of hydrochloric acid and hydrogen hydroxide) hydrometallurgical technology for processing dry sorption waste in order to extract gold and antimony. Acid treatment makes it possible to completely transfer antimony into solution, significantly improves the quality of the cake for subsequent cyanidation and gold extraction, reduces the volume of material for processing by cyanidation, and simplifies cyanidation technology. Sb2S3 activation and the possibility of lead (Pb(NO3) replacement have been studied 2) a mixture of zinc and copper (ZnSO4 and CuSO4) during the flotation of antimony ores from the Khipkoshinsky deposit in the Transbaikal Region. This has made it possible to create favorable conditions for interaction with xanthogenate and hydrophobize the surface of sulfide minerals. Theoretically, hydrogen sulfide substitutes such as NaCNS, KCNS, CuCNS for use in the flotation process were evaluated.

Effect of composition of metal-containing modifiers on flotation of sulfide minerals of nonferrous heavy metals: Analysis and modeling

Effect of composition of metal-containing modifiers on flotation of sulfide minerals of nonferrous heavy metals: Analysis and modeling

Efficient adsorption of butyl xanthate by imidazolium ionic liquid modified fumed silica

Butyl xanthate (BX) usually serve as an available collector in the sulfide mineral flotation. However, it will cause severe pollution and be harmful for living beings when discharged together with beneficiation wastewater. Adsorption, affected by sorbents structure to a great extent, is an effective strategy to treat BX-bearing wastewater. In this paper, ionic liquid-loaded fumed silica (IFS) generated by a simple process was acted as an adsorbent for removing BX from aqueous solution. The ratio of IL loaded onto silicon was about 40 %. The N elements existing in the structure of adsorbents suggested IL loaded onto silica surface successfully. The specific surface area of IFS was 65.25 m2/g, which was smaller than original material. And the total pore volume decreased from 36.75 to 15.22 cm3 g−1 after modification. Different from the virginal silica barely adsorbing BX, ionic liquid (IL) loaded silica had large uptake capacity up to 202.84 mg/g under the combined effect of functional groups modified in the silica particle and the porous structure of nanomaterial. The adsorption kinetics and isotherms are followed with the pseudo-second-order model and Langmuir model, respectively. And the functional groups, charge of adsorbent were analyzed to reveal approaching mechanism. This paper provides an alternative for xanthate and other organic contaminants treatment.

Utilizing the high intensity conditioning process with a multi-bladed impeller to enhance the flotation of fine-grained galena

Due to the increasing demand for lead metal, the recovery of fine-grained galena resources is also constantly strengthening. High intensity conditioning (HIC) has been promoted as an efficient technology for enhancing the flotation of fine minerals. In this study, the effect of the HIC process with a multi-bladed impeller on the flotation of fine-grained galena was investigated at different agitation times and speeds. By analysis of flow field characteristics, adsorption experiments, and particle aggregation observation, the correlation mechanism of HIC affecting the flotation of fine-grained galena was revealed. Compared to the results without the HIC, the flotation recovery was 95.38% when the agitation speed was 2000 rpm and the agitation time was 5 min. Mechanistic studies showed that obtaining an appropriate slurry flow rate with external energy input not only benefited the dispersion of galena particles and reagents, but also enhanced their adsorption. The appropriate impeller shear stress can increase the flow rate difference of the slurry, making it easier for fine-grained galena to form clusters with appropriate particle size, compact structure, and small porosity, which promoted mineral flotation. This study will provide a basis for adjusting the floatability of fine sulfide mineral flotation from mixing operations.

Insights into the design of polymetallic ore flotation circuits, including tailing desulfurization

Flotation is a main mineral concentration technology, which is why the design of flotation circuits has been extensively studied in the literature. However, studies on flotation circuit designs for polymetallic ores have rarely been reported in the literature. On the other hand, sulfide mineral flotation tailings are potential generators of acid mine drainage (AMD), which is one of their main environmental drawbacks. Although there are several strategies to prevent and control AMD, the most proactive attitude is to generate tailings without sulfur species. This paper investigates the circuit design of polymetallic ores, including desulfurization within the flotation process. Three types of circuit designs are analyzed: fractional flotation with the generation of a bulk concentrate and its subsequent separation, fractional flotation using selective flotation, and integrated flotation with combined bulk and selective flotation. Various aspects are analyzed, such as the effect of stage recovery uncertainty on the design; the impact of the design type on costs, energy use, and environmental effects; and the influence of the design strategy applied. Additionally, the design of a Cu–Ni plant that includes desulfurization is employed to deepen the research. The results show that it is useful to design these separation systems holistically to obtain more efficient designs. It is also concluded that the most traditional way of processing polymetallic ores consumes more energy, water, and materials, resulting in a more significant environmental impact. Finally, based on these results, a methodology for designing these complex systems is outlined.

Исследования флотационной активности синтезированных собирателей на основе фторсодержащих и ацетиленовых спиртов

Introduction. The article describes the problem of low quality mineral raw materials used for gold mining and considers the possibilities of increasing the efficiency of the technological process of beneficiation of refractory ores. As a result of the study, studies were carried out to study the flotation activity of new collecting reagents and their effect on the surface of sulfide gold-bearing minerals was studied. Research methods and materials. The methods used in the work were NMR, atomic absorption and ultraviolet spectroscopy, electron microscopy, laser diffraction, ore preparation, flotation, and electrophoretic light scattering. To determine the assessment of flotation activity as heteropolar collectors for the flotation of sulfide minerals, a number of synthesized compounds were used: 2-methyl-3-butyn-2-ol, 5-dimethyl-4-methylidene-1,3-oxathiolane-2-thione and 2. 2,3,3,4,4,5,5-octafluoropentyl sodium xanthate. Sulfide ore was used to conduct laboratory flotation studies; the monomineral pyrite was used to study the sorption of the reagent and determine the zeta potential of the surface. The following reagents were also used in the experiment: activator – copper sulfate; heteropolar collectors – potassium butylxanthate (BCX), sodium dialkyldithiophosphate (BTF-1552); organic carbon depressor – beta-sulfonaphthalene formaldehyde sodium salt (oroflos D); blowing agent – methyl isobutyl carbinol (MIBC). Research results and discussion. The synthesis of collecting reagents based on various chemical reactions was carried out and their flotation activity was studied. Several promising compounds have been obtained that have shown results when used as heteropolar collectors for the flotation of sulfide minerals. The research results showed that one of the synthesized reagents, 2-methyl-3-butyn-2-ol, has a positive effect on gold recovery during flotation. This reagent increases recovery by 0.18% and 1.46% when used together with other reagents. It also increases recovery by 1.81% when used alone. This suggests that 2-methyl-3-butyn-2-ol can be used as a replacement for one of the main heteropolar collectors. Determination of UV spectra showed the sorption of 2-methyl-3-butyn-2-ol on the surface of pyrite with an amount of up to 2.388·10-5 mol/m2 depending on the initial concentration of the reagent. The results of determining the zeta potential showed that the reagents significantly reduce this indicator, indicating electrostatic interaction and adsorption of the reagents on the pyrite surface. Conclusion. Overall, this paper describes the development of new harvesting reagents and their effects on the surface of sulfide gold-bearing minerals. The results of the study suggest the possibility of using 2-methyl-3-butyn-2-ol as a replacement for one of the main collectors in flotation. Resume. The article presents the results of studies assessing the flotation activity of a number of synthesized compounds: 2-methyl-3-butyn-2-ol, 5-dimethyl-4-methylidene-1,3-oxathiolan-2-thione and 2,2,3,3 ,4,4,5,5-octafluoropentyl sodium xanthate. The best technological performance during laboratory flotation was found for the reagent 2-methyl-3-butyn-2-ol; the extraction of gold into the concentrate was 84.68%, the content was 10.2%. For this compound, its amount adsorbed on the surface of pyrite was determined to be up to 2.388·10-5 mol/m2 in experiments. A low negative charge on the surface of pyrite was revealed after mixing it with reagent solutions when determining the zeta potential. Suggestions for practical applications and directions for future research. The results of the work can be recommended for further pilot tests with the subsequent development of technology for the industrial production of these reagents in the case of improved technological indicators of flotation gold extraction.

Perspectives on the use of biochar in the valorization of mining wastes from sulfide minerals flotation: Recovery of metals and effects on toxicity

This study aims to evaluate the use of two biochars obtained by pyrolysis of sugarcane-bagasse and compare it with commercial activated carbons as catalysts for the recovery of metals from one mining waste from sulfide minerals flotation (MW). It is also intended to determine the influence of carbon materials on the toxicity of the final residues. Leaching tests were performed in 250 mL erlenmeyer flasks using plates with magnetic stirrers during 24 h, at 90 °C and a stirring speed of 350 rpm. For each test, 5 g of MW were mixed with carbon material in two ratios of MW/carbon material (1/0.1 and 1/0.2 wt/wt) and 100 mL of leaching agent (H2SO4 solution at pH = 0.8–0.9 and 5 gL−1 of Fe3+). The experimental results showed that the addition of biochar and activated carbon enhances the recovery of Cu and Zn. The use of commercial activated carbons in ratios of 1/0.1 and 1/0.2 MW/carbon material leads to the extraction of more than 91 % of Cu and 97 % of Zn, after 24 h of leaching. For biochars, the highest recovery values of Cu (82.9 %) and Zn (98.1 %) were achieved with biochar prepared at 750 °C and used in the ratio of 1/0.2. However, the addition of carbon materials does not improve the recovery of Co. The presence of carbon materials decreased the electrical conductivity and pH of the final residue. The leaching of samples MW + W35 (1/0.1) and MW + BC550 (1/0.1) leads to a germination index higher than 90 %. For two biochars, all samples showed non-phytotoxicity.

A green flotation collector towards efficient separation of chalcopyrite from pyrite with high selectivity

Xanthate, due to its superior collecting ability, is widely used in the field of minerals flotation. However, its weak selectivity and foul odor make it difficult to separate chalcopyrite from pyrite and reduce pollution. This work provides an efficient and odorless collector to selectively separate chalcopyrite from pyrite. Amide group was introduced into xanthate and an efficient and odorless xanthate derivative was designed, potassium O-(6-(butylamino)-6-oxohexyl) xanthate (PBAHX), to selectively separate chalcopyrite and pyrite. The flotation results showed that PBAHX exhibited superior selectivity, the separation index of PBAHX for chalcopyrite and pyrite is 2.65 while that of the traditional sodium isobutyl xanthate (SIBX) is 1.29, and enabled the separation of chalcopyrite and pyrite. The electronic-nose detection demonstrated that the odor intensity of PBAHX was lower than SIBX. Subsequently, “horizontal” adsorption model was formed for the adsorption of PBAHX on pyrite to explain the differences in flotation results. Besides, hydrogen bonds may be produced between collector molecules to dampen odor. PBAHX shows potential as a green and effective alternative to sulfide mineral flotation.

New insights into mechanisms of pyrite flotation enhancement by hydrodynamic cavitation nanobubbles

Although it is well known that cavitation nanobubbles significantly improve the flotation performance of mineral particles, the effects of nanobubbles on sulfide mineral surface oxidation and reagent adsorption have not been studied sufficiently. The present study was conducted to investigate the effects of cavitation nanobubbles on oxidation characteristics and xanthate adsorption behavior of pyrite surface by means of micro-flotation tests, UV–visible spectroscopy, X-ray photoelectron spectroscopy (XPS) and zeta potential measurements. The experimental results have shown that the presence of cavitation nanobubbles in the pyrite flotation pulp enhanced the floatability and flotation kinetics of pyrite and improved the adsorption capacity of sodium amyl xanthate (SAX) and dixanthogen on pyrite surface as a result of stronger hydrophobic interactions. The reactive species generated by nanobubbles were identified to be dissolved oxygen (DO) and ·OH by quenching experiments. In addition, the zeta potential measurements have demonstrated that the rapid formation of cavitation nanobubbles on pyrite reduced the oxidation of pyrite surface. The results from this study have revealed important new mechanisms of cavitation nanobubbles enhancing sulfide mineral flotation from the perspectives of mineral surface oxidation and collector adsorption capacity and their synergistic effects that originate from unique characteristics of cavitation nanobubbles.

The depression mechanism on pyrite in a low-alkaline system with combined depressants: Experiment, HSC, DFT and ToF–SIMS studies

Depression of pyrite in a low-alkaline system has sparked soaring interests for the multi-metal sulfide minerals flotation recently. This study investigates effects of combined depressants (Ca(ClO)<sub>2</sub> and CaO) on pyrite flotation with butyl xanthate (KBX). Micro-flotation experiments indicate that the addition of 200 mg/L combined depressants (a mass ratios of CaO to Ca(ClO)<sub>2</sub> of 2:3) and 1.0×10<sup>−3</sup> mol/L KBX at pH 9.5 can effectively depresses the flotation of pyrite, and a minimum pyrite recovery rate of 12.5% is obtained. Basic thermodynamic evaluation results confirm the participation of Ca(ClO)2 significantly decrease the negative Gibbs free energies of pyrite oxidation reaction. Besides, the calcium species (Ca(OH)<sub>2</sub>, Ca<sup>2+</sup> and Ca(Cl)<sub>2</sub>) will spontaneously transform into CaCO<sub>3</sub>, and it is the ultimate dominant calcium species in the CO<sub>3</sub><sup>2-</sup> system. Density functional theory (DFT) results indicate that CaCO<sub>3</sub> can chemically adsorb onto the pyrite surface with an adsorption energy of –671.13 kJ/mol. The O1 and Ca atoms mainly contribute to the bonding process and are responsible for the stable adsorption of CaCO<sub>3</sub>. ToF-SIMS results provide strong evidence that the combined depressants increase the amount of hydrophilic species and decrease dixanthogen adsorption onto the pyrite surface. The thickness of the whole formed hydrophilic species is approximately 50 nm. Semiquantitative amounts of hydrophilic species follow the order of hydroxy calcium>iron carbonyl>calcium carbonate. Overall, hydrophilic species repulse adsorption of dixanthogen and significantly reduce the flotation performance of pyrite.

Application of inverse gas chromatography to bench scale flotation of sulphide ore

ABSTRACT Inverse gas chromatography (IGC) is receiving increasing attention due to its high precision and ease of application in determining various characteristics of a wide variety of materials of different shapes and morphologies. This study details an experimental investigation into the application of the IGC technique to the flotation of sulphide minerals. Rather than giving the surface energy, the trends associated with the various concentrates and tailings are described. Bench-scale flotation experiments of a nickel-copper sulphide ore were conducted in a Denver flotation cell. The IGC analyses were carried out on the timed concentrates, as well as on the final tailings, in order to evaluate the correlation between surface energetics and the flotation response of the ore particles. The results indicated that the floatability of the concentrates was directly related to the surface energy of the particles. Both dispersive and specific components of surface free energy increased by increasing the necessary time for the particles to be floated, which was consistent with the obtained values for the work of adhesion to water. However, the flotation response of the tailings was not consistent with the expectations from the surface energy value. The significance of the sample composition, particle size distribution, and their consequences in surface energy-flotation response relationship were also observed.

Activation of quartz flotation by Cu2+, Ni2+ in the sodium ethylxanthogenate (EX) system

During the flotation of metal sulfide minerals, due to the interference of unavoidable ions, the quartz also partially floats in some cases. The research on the mechanism of quartz being activated and floating up is still insufficient. In this study, the influence of unavoidable ions Cu2+ and Ni2+ on the floatation of quartz was studied by micro-flotation experiments, adsorption detection, zeta potential measurement, solution composition calculation, infrared spectroscopy (FT-IR) analysis, X-ray photoelectron spectroscopy (XPS) analysis, and atomic force microscopy (AFM) observation. This provides a theoretical reference for further understanding the mechanism of sodium ethylxanthogenate and quartz surface, as well as the development of new quartz depressant. The results of flotation showed that after activation by Cu2+(1×10-4 mol/L) and Ni2+(5×10-5 mol/L), quartz was captured by sodium ethylxanthogenate (EX:1.4×10-4 mol/L) under alkaline conditions, while the best recoveries were 80% and 43%, respectively. The results of adsorption measurement and Zeta potential measurement showed that the precipitation rate of Cu2+ was greater than that of Ni2+ under alkaline conditions. In addition, both Cu2+ and Ni2+ were electrostatically adsorbed on the quartz surface and changed the Zeta potential of quartz. The solution composition calculation further showed that Cu(OH)+, Cu(OH)2(s), and Ni(OH)+, Ni(OH)2(s) were the main components in the solution under alkaline conditions. FT-IR analysis, XPS analysis, and AFM observations revealed that Cu species and Ni species were adsorbed on O atoms on the quartz surface, providing active sites for EX adsorption.

Advanced collector-free flotation of typical sulfide minerals using a novel heterocyclic depressant

Molybdenite is the main source of molybdenum which is widely used in industry. Due to its excellent natural floatability, it is urgent to develop selective depressants, which is the key to achieve the flotation separation of molybdenite from other sulfide minerals such as chalcopyrite, galena and pyrite. In this work, a thiazole depressant 2,4-thiazolidinedione (CSD7) is developed based on the selective chelation of heteroatoms to Cu2+, Pb2+ and Fe2+. The flotation tests of single and mixed minerals confirm that CSD7 can achieve an efficient separation of molybdenite from chalcopyrite, galena and pyrite in a collector-free flotation system. Under the concentration of CSD7 with 5 × 10−4 mol/L and pulp pH of 7, the grades of molybdenite concentrate are higher than 50%, and the corresponding recoveries are higher than 80%. The Fourier Transform Infrared Spectroscopy (FT-IR) and electrochemical tests indicate that CSD7 can adsorb onto the surfaces of chalcopyrite, galena and pyrite, but can hardly adsorb onto the surface of molybdenite. CSD7 chelates with Cu2+, Pb2+ and Fe2+ on the surfaces of chalcopyrite, galena and pyrite through the four-membered ring formed by the bonds of −S−M2+ and =O−M2+. Moreover, the contact angle measurements also manifest that CSD7 selectively strengthen the hydrophilicity of the surfaces of chalcopyrite, galena and pyrite rather than molybdenite. The simply synthesized, low-cost and green CSD7 has a great potential in the separation of molybdenum from other sulfide minerals in flotation plant.

Molecular interactions between thiolate collectors and non-ionic frothers in mixed monolayers at the air–water interface using a regular solution theory approach

Dithiophosphate (DTP) is classically used as a co-collector in sulfide mineral flotation to promote hydrophobicity on target particles. Notwithstanding the improved recoveries observed, DTP does not appear to behave as a conventional collector, since it has been shown that it does not adsorb onto certain sulfide mineral surfaces. Recent investigations have nevertheless indicated that when DTP does not adsorb onto the sulfide mineral surface, the presence of a non-ionic frother, such as an alcohol or polyglycol ether, activates a synergistic increase in pulp phase mineral recovery and bubble-particle attachment probability. In this paper, a regular solution approach is used to quantify the interactions between thiolate collectors and non-ionic frothers at the air–water interface, where no other methodology could demonstrate such interactions. Attractive interactions were observed between DTP and frothers, whereas no such interaction was observed between xanthate, a standard base metal sulfide collector. It is suggested that this is the first time that such interactions have been shown for the DTP-frother system and this may represent an important contribution to our understanding of the role of DTP reagents in the flotation process.

Nano-microbubbles and feed size interaction in lead and zinc sulfide minerals flotation

Low recovery of the fully or highly liberated fine particles to the concentrate, and consequently, valuable loss to the tailing is a challenge of flotation industry, majorly due to the low probability of bubble-particle collision. Bubble size significantly affects the efficiency of collision. In this study, the effect of nano-micro bubbles and feed particle size (d80s in the range of 80 to 12 µm), on the lead and zinc minerals recovery, separation efficiency, and valuables loss to the tailings were investigated. The presence of NMBW simultaneously improved concentrate grade and metal recovery to the individual concentrates. Over 14.5% improvements in either lead and zinc recoveries, 0.95% and 2.22% enhancement in lead and zinc concentrate grades, and 15.6% and 8.9% increases in lead and zinc separation efficiencies were, respectively, gained. The application of NMBW, however, reduced metal loss to the tailings, more evident for the finest feed (order based on d100: -38>-75>-106 µm). By increasing NMBW content to 100% of the pulp water, the metal loss to the tailing for the dry-ground feed with the d80 of 12 µm, was decreased from 80.4% to 65.9% in the lead flotation, and from 54.9% to 43.2% in the zinc flotation.