Seaweed glue as a novel polymer depressant for the selective separation of chalcopyrite and galena

The use of seaweed glue(SEG) as a novel polymer depressant for the separation of chalcopyrite from galena with butyl xanthate(BX) as a collector was studied through microflotation experiments. Contact angle, adsorption, dynamic potential, and infrared spectral analyses were conducted to clarify the mechanism underlying the action of SEG on minerals. The results of microflotation experiments indicated that chalcopyrite could be selectively separated from galena by using a SEG depressant concentration of 15 mg·L~(-1), BX concentration of 10 mg·L~(-1), and methyl isobutyl carbinol concentration of 8.5 mg·L~(-1) at pH 8.0. A Cu concentrate with a grading of 23.68 wt% was obtained at a recovery rate of 81.52% from mixed minerals with 8.29 wt% Cu content. Contact angle analysis showed that the effect of SEG on the wettability of galena was stronger than that on the wettability of chalcopyrite. Adsorption, zeta potential, and FT-IR spectral analyses revealed that SEG and BX were coadsorbed on the surfaces of galena. SEG depressed galena by covering xanthate ions in the functional groups of –COO and mainly underwent weak physisorption on chalcopyrite. These mechanisms account for the ability of SEG to depress galena effectively while enabling chalcopyrite flotation.     

The selective effect of food-grade guar gum on chalcopyrite-monoclinic pyrrhotite separation using mixed aerofloat (CSU11) as collector

The flotation separation of chalcopyrite from monoclinic pyrrhotite using food-grade guar gum (FGG) as a depressant was studied through flotation tests, kinetic studies, dynamic potential measurements, adsorption experiments, and infrared spectral analyses. The microflotation results showed that the flotation separation of chalcopyrite from monoclinic pyrrhotite could not be realized by adding mixed aerofloat (CSU11) alone. The depressant FGG exhibited a selective depression effect on monoclinic pyrrhotite by controlling the pulp pH range from 5.0 to 6.0, with a maximum floatability variation of 79.36% in the presence of CSU11. The flotation kinetics, zeta-potential, adsorption, and infrared spectroscopy studies revealed that the FGG could absorb more strongly on the surface of monoclinic pyrrhotite than on the surface of chalcopyrite. In addition, the results revealed that the interaction of FGG with the monoclinic pyrrhotite surface was governed primarily by strong chemisorption, whereas FGG mainly bonded to chalcopyrite through hydrogen bonding. This difference was responsible for the excellent depression selectivity of FGG toward monoclinic pyrrhotite flotation and weak depression effect toward chalcopyrite flotation.     

Selective depression behavior of guar gum on talc-type scheelite flotation

The depression behavior and mechanism of guar gum on talc-type scheelite flotation were systematically investigated by flotation experiments, adsorption tests, zeta-potential measurements, and infrared spectroscopic analyses. The flotation results for monominerals, mixed minerals, and actual mineral samples indicated that guar gum exhibited much higher selective depression for talc than for scheelite. Bench-scale closed-circuit tests showed that a tungsten concentrate with a WO₃ grade of 51.43% and a WO₃ recovery of 76.18% was obtained. Adsorption tests, zeta-potential measurements, and infrared spectral analyses confirmed that guar gum absorbed more strongly onto the talc surface than onto the scheelite surface because of chemisorption between guar gum and talc. This chemisorption is responsible for the guar gum's highly selective depression for talc and small depression for scheelite. The flotation results provide technical support for talc-type scheelite flotation.     

Solution chemistry of carbonate minerals and its effects on the flotation of hematite with sodium oleate

The effects of carbonate minerals (dolomite and siderite) on the flotation of hematite using sodium oleate as a collector were investigated through flotation tests, supplemented by dissolution measurements, solution chemistry calculations, zeta-potential measurements, Fourier transform infrared (FTIR) spectroscopic studies, and X-ray photoelectron spectroscopy (XPS) analyses. The results of flotation tests show that the presence of siderite or dolomite reduced the recovery of hematite and that the inhibiting effects of dolomite were stronger. Dissolution measurements, solution chemistry calculations, and flotation tests confirmed that both the cations (Ca2+ and Mg2+) and CO₃2- ions dissolved from dolomite depressed hematite flotation, whereas only the CO₃2- ions dissolved from siderite were responsible for hematite depression. The zeta-potential, FTIR spectroscopic, and XPS analyses indicated that Ca2+, Mg2+, and CO₃2- (HCO₃-) could adsorb onto the hematite surface, thereby hindering the adsorption of sodium oleate, which was the main reason for the inhibiting effects of carbonate minerals on hematite flotation.     

Ferric ion-triggered surface oxidation of galena for efficient chalcopyrite–galena separation

The efficient separation of chalcopyrite(CuFeS₂) and galena(PbS) is essential for optimal resource utilization. However, finding a selective depressant that is environmentally friendly and cost effective remains a challenge. Through various techniques, such as microflotation tests, Fourier transform infrared spectroscopy, scanning electron microscopy(SEM) observation, X-ray photoelectron spectroscopy(XPS), and Raman spectroscopy measurements, this study explored the use of ferric ions...     

Effectiveness of sodium silicate as gangue depressants in iron ore slimes flotation

The recovery of iron from the screw classifier overflow slimes by direct flotation was studied. The relative effectiveness of sodium silicates with different silica-to-soda mole ratios as depressants for silica and silicate bearing minerals was investigated. Silica-to-soda mole ratio and silicate dosage were found to have significant effect on the separation efficiency. The results show that an increase of Fe content in the concentrate is observed with concomitant reduction in SiO₂ and Al₂O₃ levels when a particular type of sodium silicate at a proper dosage is used. The concentrate of 58.89wt% Fe, 4.68wt% SiO₂, and 5.28wt% Al₂O₃ with the weight recovery of 38.74% and the metal recovery of 41.13% can be obtained from the iron ore slimes with 54.44wt% Fe, 6.72wt% SiO₂, and 6.80wt% Al₂O₃, when Na₂SiO₃ with a silica-to-soda mole ratio of 2.19 is used as a depressant at a feed rate of 0.2 kg/t.     

Use of centrifugal-gravity concentration for rejection of talc and recovery improvement in base-metal flotation

The possibility of using a centrifugal-gravity concentrator to reject Mg-bearing minerals and minimize metal losses in the flotation of base metals was evaluated. Sample characterization, batch scoping tests, pilot-scale tests, and regrind-flotation tests were conducted on a Ni flotation tailings stream. Batch tests revealed that the Mg grade decreased dramatically in the concentrate products. Pilot-scale testing of a continuous centrifugal concentrator (Knelson CVD6) on the flotation tailings revealed that a concentrate with a low mass yield, low Mg content, and high Ni upgrade ratio could be achieved. Under optimum conditions, a concentrate at 6.7% mass yield was obtained with 0.85% Ni grade at 12.9% Ni recovery and with a low Mg distribution (1.7%). Size partition curves demonstrated that the CVD also operated as a size classifier, enhancing the rejection of talc fines. Overall, the CVD was capable of rejecting Mg-bearing minerals. Moreover, an opportunity exists for the novel use of centrifugal-gravity concentration for scavenging flotation tailings and/or after comminution to minimize amount of Mg-bearing minerals reporting to flotation.     

Process mineralogy as a key factor affecting the flotation kinetics of copper sulfide minerals

The aim of this study is to apply process mineralogy as a practical tool for further understanding and predicting the flotation kinetics of the copper sulfide minerals. The minerals' composition and association, grain distribution, and liberation within the ore samples were analyzed in the feed, concentrate, and the tailings of the flotation processes with two pulp densities of 25 wt% and 30 wt%. The major copper-bearing minerals identified by microscopic analysis of the concentrate samples included chalcopyrite(56.2 wt%), chalcocite(29.1 wt%),covellite(6.4 wt%), and bornite(4.7 wt%). Pyrite was the main sulfide gangue mineral(3.6 wt%) in the concentrates. A 95% degree of liberation with d_(80) 80 μm was obtained for chalcopyrite as the main copper mineral in the ore sample. The recovery rate and the grade in the concentrates were enhanced with increasing chalcopyrite particle size. Chalcopyrite particles with a d_(80) of approximately 100 μm were recovered at the early stages of the flotation process. The kinetic studies showed that the kinetic second-order rectangular distribution model perfectly fit the flotation test data. Characterization of the kinetic parameters indicated that the optimum granulation distribution range for achieving a maximum flotation rate for chalcopyrite particles was between the sizes 50 and 55 μm.     

Comprehensive recovery of gold and base-metal sulfide minerals from a low-grade refractory ore

The comprehensive recovery of small amounts of valuable minerals such as gold and base-metal sulfide minerals from a low-grade refractory ore was investigated. The following treatment strategy was applied to a sample of this ore: gold flotation–gold concentrate leaching–lead and zinc flotation from the gold concentrate leaching residue. Closed-circuit trials of gold flotation yielded a gold concentrate that assayed at 40.23 g·t-1 Au with a recovery of 86.25%. The gold concentrate leaching rate was 98.76%. Two variants of lead-zinc flotation from the residue—preferential flotation of lead and zinc and bulk flotation of lead and zinc—were tested using the middling processing method. Foam from the reflotation was returned to the lead rougher flotation or lead–zinc bulk flotation, whereas middlings from reflotation were discarded. Sulfur concentrate was a byproduct. The combined strategy of flotation, leaching, and flotation is recommended for the treatment of this kind of ore.     

利用自动解离分析方法研究硫化铜尾矿的矿物学特征:谦比希铜矿尾矿的实例研究

As ore grades constantly decline, more copper tailings, which still contain a considerable amount of unrecovered copper, are expected to be produced as a byproduct of froth flotation. This research reveals the occurrence mechanism of copper minerals in typical copper sulfide tailings using quantitative mineral liberation analysis (MLA) integrated with scanning electron microscopy–energy dispersive spectroscopy (SEM–EDS). A comprehensive mineralogical characterization was carried out, and the results showed that almost all copper minerals were highly disseminated within coarse gangue particles, except for 9.2wt% chalcopyrite that occurred in the 160–180 μm size fraction. The predominant copper-bearing mineral was chalcopyrite, which was closely intergrown with orthoclase and muscovite rather than quartz. The flotation tailings sample still contained 3.28wt% liberated chalcopyrite and 3.13wt% liberated bornite because of their extremely fine granularity. The SEM–EDS analysis further demonstrated that copper minerals mainly occurred as fine dispersed and fully enclosed structures in gangue minerals. The information obtained from this research could offer useful references for recovering residual copper from flotation tailings.     

Dual effects of sodium sulfide on the flotation behavior of chalcopyrite: I. Effect of pulp potential

This study explores the flotation behavior of chalcopyrite in the presence of different concentrations of sodium sulfide (Na₂S·9H₂O) at pH 12 under controlled potential conditions. It was observed that the flotation of chalcopyrite is not depressed completely when the pulp potential is low, even at high concentrations of sodium sulfide, i.e., 10?1–10?2 mol/L. However, a partial and controlled oxidation of pulp does enhance the effectiveness of sodium sulfide on the depression of chalcopyrite. Characterization of the chalcopyrite particle surface by X-ray photoelectron spectroscopy allowed the identification of hydrophilic and hydrophobic surface species, which are responsible for the depression and flotation of chalcopyrite. Changes in pulp potential were found to be an effective float controlling parameter, by which Na2S can be used to initiate or depress the flotation behavior of chalcopyrite.     

Improved flotation performance of hematite fines using citric acid as a dispersant

In this study, citric acid was used as a dispersant to improve the flotation performance of hematite fines. The effect and mechanism of citric acid on the reverse flotation of hematite fines were investigated by flotation tests, sedimentation experiments, scanning electron microscopy (SEM), zeta-potential measurements, and X-ray photoelectron spectroscopy (XPS). The results of SEM analysis and flotation tests reveal that a strong heterocoagulation in the form of slime coating or coagulation in hematite fine slurry affects the beneficiation of hematite ores by froth flotation. The addition of a small amount of citric acid (less than 300 g/t) favorably affects the reverse flotation of hematite fines by improving particle dispersion. The results of sedimentation experiments, zeta-potential measurements, and XPS measurements demonstrate that citric acid adsorbs onto hematite and quartz surfaces via hydrogen bonding, thereby reducing the zeta potentials of mineral surfaces, strengthening the electrical double-layer repulsion between mineral particles, and dispersing the pulp particles.     

Mineralogical characterization of copper sulfide tailings using automated mineral liberation analysis: A case study of the Chambishi Copper Mine tailings

As ore grades constantly decline, more copper tailings, which still contain a considerable amount of unrecovered copper, are expected to be produced as a byproduct of froth flotation. This research reveals the occurrence mechanism of copper minerals in typical copper sulfide tailings using quantitative mineral liberation analysis(MLA) integrated with scanning electron microscopy–energy dispersive spectroscopy(SEM–EDS). A comprehensive mineralogical characterization was carried out, and the results showed that almost all copper minerals were highly disseminated within coarse gangue particles, except for 9.2 wt% chalcopyrite that occurred in the 160–180 μm size fraction. The predominant copper-bearing mineral was chalcopyrite, which was closely intergrown with orthoclase and muscovite rather than quartz. The flotation tailings sample still contained 3.28 wt% liberated chalcopyrite and 3.13 wt% liberated bornite because of their extremely fine granularity.The SEM–EDS analysis further demonstrated that copper minerals mainly occurred as fine dispersed and fully enclosed structures in gangue minerals. The information obtained from this research could offer useful references for recovering residual copper from flotation tailings.     

Extracting Gold Fines from Pyrite Slag by Hydrophobic Flocculation Flotation

In order to extract gold fines from pyrite slag, the hydrophobic flocculation flotation (HFF) was studied and compared with conventional flotation. The main parameters of HFF such as flotation reagent, surfactant dosage. the duration of agitation and the amount of non-polar oil were investigated, and the effect of non-polar oil was analyzed particularly. It is demonstrated that the HFF is better for gold extraction from pyrite slag than the conventional flotation, and the non-polar oil and the intensive agitation are the key to improve the Au contents and recovery in the concentrate. HFF was used to treat the pyrite slag from the west of Jiangxi province, whose gold contents are 2.949/t and the size of which are smaller than 10μm. The gold concentrate with gold grade 126.3 g/t and 51.35% recovery was obtained. Thus, a new method of extracting gold fines from pyrite slag is developed using HFF.     

Effect of reverse flotation on magnetic separation concentrates

Reverse flotation studies on magnetite samples have revealed that the use of starch as a depressant of Fe-oxides has a hydrophilic effect on the surface of Fe-bearing silicates and significantly decreases Fe in the silica-rich stream when used in combination with an amine (Lilaflot D817M). In this study, the effect of reverse flotation on the optimization of products obtained from magnetic separation was investigated. Two different magnetic samples, zones 1 and 2, were milled to <75 μm and then subjected to low intensity magnetic separation (LIMS). The LIMS test conducted on the <75 μm shown an upgrade of 46.40wt% Fe, 28.40wt% SiO₂ and 2.61wt% MnO for zone 1 and 47.60wt% Fe, 29.17wt% SiO₂ and 0.50wt% MnO for zone 2. Further milling of the ore to <25 μm resulted in a higher magnetic-rich product after magnetic separation. Reverse flotation tests were conducted on the agitated magnetic concentrate feed, and the result shows a significant upgrade of Fe compared to that obtained from the non-agitated feed. Iron concentrations greater than 69%, and SiO₂ concentrations less than 2% with overall magnetite recoveries greater than 67% and 71% were obtained for zones 1 and 2, respectively.     

Semi-industrial experimental study on bauxite separation using a cell–column integration process

The cyclonic-static micro-bubble flotation column (FCSMC) is a highly efficient mineral processing equipment. In this study, a cell-column (FCSMC) integration process was investigated for the separation of bauxite and its feasibility was analyzed on a theoretical basis. The properties of low-grade bauxite ore from Henan Province, China were analyzed. Parameters such as reagent dosage, scraping bubble time, and pressure of the circulating pump during the sorting process were investigated and optimized to improve the flotation efficiency. On the basis of these parameters, continuous separation experiments were conducted. Bauxite concentrate with an aluminum-to-silicon (A/S) mass ratio of 6.37 and a 77.63wt% recovery rate were achieved via a flow sheet consisting of “fast flotation using a flotation cell, one roughing flotation and one cleaning flotation using flotation columns”. Compared with the full-flotation-cells process, the cell–column integration process resulted in an increase of the A/S ratio by 0.41 and the recovery rate by 17.58wt%. Cell–column integration separation technology represents a new approach for the separation of middle-to-low-grade bauxite ore.     

Effect of Flotation Reagents on the Cake Moisture of Copper Concentrate

The effect of reagents used in separating chalcopyrite from pyrite on the cake moisture of the copper concentrate at Daye Iron Mine Mineral Processing Plant was investigated. The results showed that the dosage of lime used for depressing pyrite was the main factor that increased the filter cake moisture of copper concentrate. With increasing the dosage of lime, the cake moisture of copper concentrate increased sharply. The cause was concluded to be the addition of lime to the pulp, which resulted in the formation of floc and a high pH value. The collector Z-200#, used for collecting chalcopyrite, had, as well, an adverse effect on the cake moisture of copper concentrate, but its effect was inferior in respect to that of lime. The cake moisture of copper concentrate can be decreased by changing the method with which lime is added and the pH value of pulp is regulated. The experiment results showed that the sulfuric acid was the best regulator. When the clarified liquor of lime was used as a depressant and the pH value of the pulp was regulated to 6.5-7.0 by adding sulfuric acid, the cake moisture of copper concentrate was reduced from 15.49% to 13.13%. The examination of chalcopyrite surface by using ESCA (Electron Spectroscopy for Chemical Analysis) showed that calcium sulfate and iron hydroxide had formed on the surface of chalcopyrite when lime was added to the pulp. The formation of calcium sulfate and iron hydroxide on its surface increased the hydrophilicity of chalcopyrite so that its cake moisture increased. The addition of sulfuric acid to the pulp not only removed the calcium sulfate, but also reduced the concentration of iron hydroxide on the surface of chalcopyrite so that the cake moisture of copper concentrate was decreased.     

Improved hemimorphite flotation using xanthate as a collector with S(Ⅱ) and Pb(Ⅱ) activation

The flotation of hemimorphite using the S(Ⅱ)–Pb(Ⅱ)–xanthate process, which includes sulfidization with sodium sulfide, activation by lead cations, and subsequent flotation with xanthate, was investigated. The flotation results indicated that hemimorphite floats when the S(Ⅱ)–Pb(Ⅱ)–xanthate process is used; a maximum recovery of approximately 90% was obtained. Zeta-potential, contact-angle, scanning electron microscopy–energy-dispersive spectrometry (SEM–EDS), and diffuse-reflectance infrared Fourier transform spectroscopy (DRIFTS) measurements were used to characterize the activation products on the hemimorphite surface and their subsequent interaction with sodium butyl xanthate (SBX). The results showed that a ZnS coating formed on the hemimorphite surface after the sample was conditioned in an Na₂S solution. However, the formation of a ZnS coating on the hemimorphite surface did not improve hemimorphite flotation. With the subsequent addition of lead cations, PbS species formed on the mineral surface. The formation of the PbS species on the surface of hemimorphite significantly increased the adsorption capacity of SBX, forming lead xanthate (referred to as chemical adsorption) and leading to a substantial improvement in hemimorphite flotation. Our results indicate that the addition of lead cations is a critical step in the successful flotation of hemimorphite using the sulfidization–lead ion activation–xanthate process.     

Slime coating of kaolinite on chalcopyrite in saline water flotation

In saline water flotation, the salinity can cause a distinguishable slime coating of clay minerals on chalcopyrite particles through its effect on their electrical double layers in aqueous solutions. In this work, kaolinite was used as a representative clay mineral for studying slime coating during chalcopyrite flotation. The flotation of chalcopyrite in the presence and absence of kaolinite in tap water, seawater, and gypsum-saturated water and the stability of chalcopyrite and kaolinite particles in slurries are presented. Zeta-potential distributions and scanning electron microscopy images were used to characterize and explain the different slime coating degrees and the different flotation performances. Kaolinite particles induced slime coating on chalcopyrite surfaces and reduced chalcopyrite floatability to the greatest extent when the pH value was in the alkaline range. At 0.24wt% of kaolinite, the chalcopyrite floatability was depressed by more than 10% at alkaline pH levels in tap water. Salinity in seawater and gypsum-saturated water compressed the electrical double layers and resulted in extensive slime coating.     

Aerosol flotation of low-grade refractory molybdenum ores

The characteristics of aerosol flotation, which include the effect of the concentration and particle size of kerosene aerosol on the molybdenum (Mo) flotation index and the effect of kerosene aerosol dosing method on the kerosene dosage and flotation time, were studied in the flotation of low-grade refractory molybdenum ores using kerosene aerosol. The results revealed that the particle size and concentration of kerosene aerosol had little effect on the Mo grade but had significant effect on the Mo recovery. A smaller particle size and a lower concentration of kerosene aerosol were beneficial to the Mo aerosol flotation. For the received Mo ore samples, the optimized particle size of kerosene aerosol was 0.3-2 μm and the optimized aerosol concentration was 14 mg/L. The compressed air atomizer had a more uniform distribution of aerosol particles than the ultrasonic atomizer, and the aerosol concentration was controlled easily, so the compressed air atomizer was more suitable for the research of aerosol flotation. Compared with conventional flotation in which kerosene was directly added into the ore pulp, the flotation time was reduced by ~30%, and the dosage was decreased by ~20% in aerosol flotation, while the Mo flotation index was similar.