Lead Flotation Research Articles

Investigate of Effective Factors on Extraction of Silver from Tailings of Lead Flotation Plant Using Thiourea Leaching

In current paper, the effect of different factors on extraction of silver from tailings of lead flotation plant using thiourea leaching was studied. According to the mineralogical studies and chemical analysis taken from the tailings dam, the representative sample taken from there contained 30 ppm Ag and the minerals such as calcite, dolomite, barite, microcline, galena etc. In this research, the effects of iron(III) sulfate concentration, thiourea concentration, size and temperature onextraction of silver from Ravanj flotation plant tailings were analyzed using statistics design of experiment and DX7 software; and it was determined that iron(III) sulfate concentration and temperature were the most effective factors on the extraction of silver. The results showed that the best recovery obtained in particle size of 75 microns, 1 kg/m3 thiourea concentration and 1 kg/m3 iron(III) sulfate concentration in the temperature of 60℃ for 2 hours.

Study on Potential-Controlled Flotation Test of a Copper-Lead-Zinc Polymetallic Sulphide Ore

The study on potential-controlled flotation test of differential flotation process was carried out in the light of the change of a certain ore properties. The test used self-developed EMZ-91, as well as conventional collectors of ethyl thio carbamate and butyl xanthate for the flotation of copper, lead and zinc ores respectively, in which the copper sulfate was used as the activator of zinc mineral. The new differential flotation process, which is using lime to regulate pulp potential, produced the copper concentrate grading 27.18% copper at 73.37% recovery, the lead concentrate grading 66.00% lead at 63.00% recovery, and the zinc concentrate grading 55.27% zinc at 87.69% recovery.

Multi-criteria decision making for collector selection in the flotation of lead–zinc sulfide ore

The results of multi-criteria decision making (MCDM) for the selection of the optimum sulfide mineral collector and dose rate in a flotation study are presented. Laboratory scale flotation tests were performed to investigate selective flotation of galena from lead–zinc sulfide ores using several different xanthates at different dose rates. The flotation results show that the different tests result in changes lead concentrate grade, lead recoveries and selectivity against zinc. According to the selection criteria to maximize lead concentrate grades and recoveries while minimizing zinc concentrate grades as different factors, decision making and final selection of the optimum collector and dose rates are not simple since there is multiple options and outcomes to consider.A technique for order preference by similarity to ideal solution (TOPSIS), as well-known MCDM method, has been used for ranking xanthates not only using technological, but also economical factors. The results obtained show that MCDM is a useful tool for ranking and selection of tested collectors in flotation because of the ability to evaluate all factors fixed in flotation tests. The final step is a developed mathematical model for selection of optimal collector.

Anglesite and silver recovery from jarosite residues through roasting and sulfidization-flotation in zinc hydrometallurgy

Hazardous jarosite residues contain abundant valuable minerals that are difficult to be recovered by traditional flotation process. This study presents a new route, roasting combined with sulfidization-flotation, for the recovery of anglesite and silver from jarosite residues of zinc hydrometallurgy. Surface appearance and elemental distribution of jarosite residues was examined by scanning electron microscopy and energy dispersive X-ray spectrometry analysis, respectively. Decomposition and transformation mechanisms of jarosite residues were illustrated by differential thermal analysis. Results showed that after roasting combined with flotation, the grade and recovery of lead were 43.89% and 66.86%, respectively, and those of silver were 1.3kg/t and 81.60%, respectively. At 600–700°C, jarosite was decomposed to release encapsulated valuable minerals such as anglesite (PbSO4) and silver mineral; silver jarosite decomposed into silver sulfate (Ag2SO4); and zinc ferrite (ZnO·Fe2O3) decomposed into zinc sulfate (ZnSO4) and hematite (Fe2O3). Bared anglesite and silver minerals were modified by sodium sulfide and easily collected by flotation collectors. This study demonstrates that the combination of roasting and sulfidization-flotation provides a promising process for the recovery of zinc, lead, and silver from jarosite residues of zinc hydrometallurgy.

Experimental Research on Rough Flotation of Lead in Bainiuchang Lead-Zinc Mines

Containing Pb (1.09% ) and Zn (1.97%),the lead-zinc sulphide ore of Bainiuchang mine is a refractory resource in Yunnan Mengzi. The flotation performance of lead is ineffective in plant practice. This research is carried out to enhance the rough flotation performance of lead. With the dosage of lime and zinc sulfate as depressors, sodium hexametaphosphate as dispersant, and ethyl thio carbamate and MA as combined collector, the grinding fineness of 85% , a lead concentrate with lead grade of 17.68%, lead recovery rate of 75.59%, and zinc grade of 7.16% was obtained. The performance of rough flotation of lead is improved.

Study of Lead Flotation Test in Oxide Lead-Zinc Ore

This paper studied an technological process of recovery of Pb and Zn from a difficult to choose high oxidation Lead –Zinc deposit. By mineralogy process study shown, the run-of-mine grade of Lead was 4.09%, with the oxidation rate of 88.26%. The optimal grinding fineness was about to 85.68%. Use the best reagent combination which obtained by condition test, the final results was that the grade of Lead concentrate was 35.6% and the recovery rate was 78.73%, included silver 564 g/t; the grade of the Zinc concentrate was 42.50%, and the recovery rate is 83.65%.

Arsenic content in the lead flotation concentrate in dependence on NaCN dosage

Flotation properties of arsenopyrite are similar to the flotation properties of pyrite. Specifically, in the alkaline conditions (pH value of pulp 7-11) and in the presence of xanthates as collectors, arsenopyrite can easily pass into concentrate. Therefore, if the ore contains arsenopyrite as supporting component, its flotation characteristics can affect the certain difficulties in the flotation of other minerals (Pb, Zn, Sb, W, etc.), because arsenic is considered as harmful element in almost all concentrates of metallic minerals. To the aim of obtaining high quality flotation concentrate, it is often necessary to depress arsenopyrite, which can be successfully achieved by applying of cyanides. This paper presents the results of laboratory tests, performed in order to determine the possibility of obtaining a lead concentrate with low content of arsenic. Ore sample, which was used in this flotation testing, contained a large amount of arsenopyrite (over 40% by mass). The results showed that arsenic content in the rough lead concentrate can be decreased up to ten times compared to its content in the ore, using NaCN as depressant.

Recycle of Wastewater from Lead-Zinc Sulfide Ore Flo-tation Process by Ozone/BAC Technology

Lead-zinc sulphide ore contains lead sulphide (galena), and zinc sulphide (sphalerite). In the first flotation stage, galena is rendered hydrophobic with an organic collector such as xanthate, while sphalerite is kept from floating by depressants, and in the second flotation stage, activator was used to activated zinc flotation. Since the organic regent used are different in the two flotation stage, wastewater from the second zinc flotation stage can’t be directly recycled to the first lead flotation stage. Wastewater from flotation process for concentrating lead-zinc sulphide ore often containing organic compounds such as diethyldithiocarbamate(DDTC), xanthate, terpenic oil(2# oil) and thionocarbamate esters (Z-200), are environmentally hazardous. Their removal from contaminated water and the reuse of the water is one of the main challenges facing lead-zinc sulphide ore processing plants. In this study, synthetic wastewater containing DDTC, xanthate, 2# oil and Z-200 at concentrations ranging from 21 to 42 mg/L was fed into an Ozone/Biological activated carbon (BAC) reactor. Analyses of the effluent indicated a chemical oxygen demand (COD) removal over 86.21% and Total organic carbon (TOC) removal over 90.00% were achieved under Hydraulic retention time (HRT) of 4h and O3 feeding concentration of 33.3mg/L. The effluent was further recycled to the lab scale lead concentrating process and no significant difference was found in compare with fresh water. Furthermore, lead-zinc sulphide mineral concentrating process was carried out at lab scale. The produced wastewater was treated by Ozone/BAC reactor at O3 feeding concentration of 16.7mg/L and HRT of 4h. The effluent analysis showed that TOC removal was 74.58%. This effluent was recycled to the lab scale lead-zinc sulphide mineral concentrating process and the recovery of lead was not affected. The results showed that by using Ozone/BAC technology, the lead-zinc sulphide mineral processing wastewater could be recycled.

Implementation of flotation and gravity separation, to process Changarzeh sulfide-oxide lead ore

This study aimed to explore concentration of a low grade sulfide-oxide lead-zinc sample containing 2.3% Pb, 1.91% Zn taken from Changarzeh mine, South Natanz, Iran. The effects of different parameters such as type and dosage of collector, milling retention time, dosage of sodium sulfur and its preparation time, application of sodium silicate, pH and solid content were investigated in relation to flotation efficiencies. Optimum experiment was carried out in cumulative flotation with 200g/t KAX as collector, 2000g/t Na2S, 500g/t sodium silicate, 30g/t MIBC and at pH=10. This test yielded 94% of lead recovery under optimum condition performance, and a concentrate with 70%Pb was produced through cleaning stages. Eighteen minutes was found to be the optimum time for lead flotation in laboratory scale. Application of gravity method for the production of a middle product was also considered. For shaking table the effect of table slope and water flowrate and for jig the effect of water flowrate and frequency were studied. Gravity separation by shaking table resulted in a concentrate with 46%Pb and 80% recovery, so shaking table could be proposed for production of pre-concentrate.

Multi-Objective Optimization of Some Process Parameters of Lead Flotation Using Grey Relational Analysis

This paper presents a novel effective method for the optimization of a lead flotation with multiple performance characteristics based on the grey relational analysis (GRA). The lead flotation parameters including potassium amyl xanthate (KAX) as a collector, sodium sulfide (Na2S), and slurry pH were optimized based on multiple performance characteristics. Sixteen experiments have been conducted using GRA to optimize the settings for the lead flotation parameters to generate two quality characteristics (grade and recovery of the concentrate). Analysis of the grey relational grade indicates that parameter significance and the optimal parameter combination for the lead flotation process are identified. The analytical results from two confirmation experiments using the optimal parameters confirm that the above performance characteristics in the lead flotation process can be improved effectively through this approach.

Direct selective lead, copper and zinc minerals flotation from polymetallic ore “Podvirovi”

The laboratory tests to find out the possibility of direct selective flotation of lead, copper and zinc minerals from the ore deposit “Podvirovi” were submitted to the Mineral Processing Department of the Faculty of Mining and geology, University of Belgrade. Based on thus gained results, it has been concluded that it is possible to obtain selective concentrate of lead copper and zinc of suitable quality, with satisfactory recovery.

Ion Flotation of Copper(II) and Lead(II) from Environmental Water Samples

The present study aims to develop a simple, rapid and economic procedure for copper(II) and lead(II) removal under the optimum conditions investigated. It is based on the complex formation between Cu2+ and Pb2+ ions and diphenylcarbazone (HDPC) followed by flotation with oleic acid (HOL) surfactant. The different parameters (namely: solution pH, HDPC, HOL, copper and lead concentrations, ionic strength, temperature and the presence of foreign ions) influencing the flotation process were examined. Nearly, 100% of Cu2+ and Pb2+ ions were removed from aqueous solutions at pHs 6 and 7, respectively at room temperature (~25oC). The procedure was successfully applied to recover almost copper(II) and lead(II) spiked to some natural water samples. Due to the rapid, simple and economic nature of the procedure, a flotation mechanism is suggestedfor metal removal in wastewater systems

Depression of pyrite in the flotation of high pyrite low-grade lead–zinc ore using Acidithiobacillus ferrooxidans

In this research work, selective depression of pyrite in the flotation of a low-grade lead–zinc ore containing 31% pyrite was investigated in the absence and presence of Acidithiobacillus ferrooxidans. Pyrite was significantly depressed with these bacteria using the optimum dosage of reagents. Sphalerite recovery and Zn grade in the obtained sphalerite concentrate were both enhanced by bacteria. The results of bioflotation experiments showed that bacterial depression of pyrite is very sensitive to the concentration of other flotation reagents. Least significance difference (LSD) bars were used to study the significance of the factors under study with a confidence interval of 90%. Under similar conditions, A. ferrooxidans was seen to increase the Zn grade and separation efficiency more effectively than sodium cyanide. Flotation kinetic studies confirmed a considerable decrease in the pyrite kinetic rate constant in the presence of A. ferrooxidans.

Dispersion effect on a lead–zinc sulphide ore flotation

This investigation was performed with samples from a lead–zinc sulphide deposit aiming at studying the influence of the dispersion degree of the particles in the pulp on lead and zinc flotation. Samples of ore and also of the minerals sphalerite, galena, pyrite, and dolomite were selected for the experiments. Nine types of dispersing agents and six blends among them were employed. A set of three dispersing agents was selected for the lead flotation and another set of three was chosen for zinc flotation. The criteria for the reagents selection were: high dispersion degree for galena and low for the other species, high dispersion degree for sphalerite and low for the other species, low dispersion degree for pyrite and high for the other species, and high dispersion degree for all species. Lead flotation experiments were performed under three conditions aiming at verifying the influence of the dispersing agent, of the pH, and of sodium carbonate. The zinc flotation tests were carried out at pH 10.5, modulated with lime. The use of dispersing agents in lead flotation did not improve the overall efficiency of the circuit for, despite improving the lead metallurgical recovery, they increase significantly the zinc losses in the lead concentrate. Sodium carbonate presented a low dispersion degree and did not affect the lead flotation results when compared with those achieved at natural pH and at pH 9.8 modulated with lime. Two dispersing agents were particularly effective in zinc flotation: dispersant 3223, a sodium polyacrylate, and sodium hexametaphosphate. Both reagents significantly enhanced zinc recovery without impairing the concentrate quality.

Separation of oxidized zinc minerals from tailings: Influence of flotation reagents

Lead flotation tailings of Dandy mineral processing plant in north-western Iran contain oxidized zinc minerals. Zinc recovery from the tailing was studied in the presence of different flotation reagents. Microscopic analysis showed that the major phases in the tailing were smithsonite and quartz. A sulfidization–flotation method has been used on a laboratory scale to investigate the effect of various reagents such as sodium sulfide as sulfidizing agent, primary amine as collector, dispersants and flocculants. To reach an optimum zinc recovery, different concentrations of the reagents were also studied. Smithsonite is friable and a large amount of fines and slimes produce. Usually de-sliming is used to prevent slime coating and to increase the recovery of zinc. However, analyses have shown that fine particles in the sample mostly contain zinc, thus de-sliming was not suggested. To enhance zinc recovery an effective disperssing reagent could increase the flotation selectivity of these tailings. Among the dispersants tested, sodium hexametaphosphate gave a higher zinc grade of 40.7% and 70% recovery. Comparing to the dispersants used, it seems that the flocculant is not that effective in increasing the zinc recovery, simply coagulating the fine particles.

Optimization of Pb flotation using statistical technique and quadratic programming

A three-level Box–Behnken factorial design combining with a response surface methodology (RSM) was employed for modeling and optimizing three operations parameters of the flotation on lead flotation. The flotation studies of lead ores containing 7.1% Pb were carried out in this study. The variables studied were potassium amyl xanthate (KAX) as a collector, sodium sulfide (Na 2S) and pH. Experiments were carried out using Box–Behnken factorial design. The main and interaction effects were evaluated using response surface methodology. Second order response functions were produced for both the Pb grade and recovery. These response functions were then optimized using the quadratic programming method to maximize Pb grade and recovery within the experimental range studied. The optimum composition was found to be 212 g/t KAX, 1250 g/t sodium sulfide and 9 pH to achieve the maximum Pb grade. The model prediction of 47.44% Pb at optimum conditions was higher than any value obtained in the initial experiments conducted. In the same way, the optimum composition was found to be 100 g/t KAX, 1250 g/t Na 2S and 7 pH to achieve the maximum recovery. The model prediction of 84.58% recovery at optimum conditions was also higher than any value obtained in the initial experiments conducted.

Hydrometallurgical treatment of tailings with high zinc content

Zinc exists as smithsonite and hemimorphite in the lead flotation tailings from the Dandi mineral processing plant in north western Iran. In this research, zinc-rich tailings produced in the Dandi plant were characterized mineralogically and a leaching study was carried out to assess the effect of several parameters on the kinetics of zinc dissolution. Parameters studied included: sulfuric acid concentration, reaction time, temperature and slurry density. It was found that leaching is controlled by a single rate-controlling step with an activation energy of 23.5 kJ/mol. To overcome some of the filtration problems associated with polymerization of silicic acid, lime was added as a coagulant. The optimum pH, holding time and temperature required to maximize the filtration rate were determined.

Effect of cuprous cyanide, dry and wet milling on the selective flotation of galena and sphalerite

Batch flotation tests of a lead–zinc sulphide composite ore from the Rosh Pinah Mine have been carried out at pH 8.5 in the presence of copper cyanide complexes. These copper cyanide complexes are often found in the recycled water that is used in the milling and the lead flotation circuits. Flotation results have shown that cuprous cyanide complexes can activate sphalerite. In addition, the activation and subsequent flotation of sphalerite was greater when the composite was dry milled as compared to wet milling. Surface analysis of copper(I)-activated sphalerite samples was studied by X-ray photoelectron spectroscopy (XPS). XPS results showed the presence of copper species on the surface of sphalerite after activation with cuprous cyanide complexes. The copper species could be removed from the surface of sphalerite after treatment with sodium cyanide. This explains, amongst others reasons, the high cyanide requirement at Rosh Pinah Mine for the efficient depression of sphalerite in the lead flotation circuit.

QSAR modeling of flotation collectors using principal components extracted from topological indices.

Several topological indices were calculated for substituted-cupferrons that were tested as collectors for the froth flotation of uranium. The principal component analysis (PCA) was used for data reduction. Seven principal components (PC) were found to account for 98.6% of the variance among the computed indices. The principal components thus extracted were used in stepwise regression analyses to construct regression models for the prediction of separation efficiencies (Es) of the collectors. A two-parameter model with a correlation coefficient of 0.889 and a three-parameter model with a correlation coefficient of 0.913 were formed. PCs were found to be better than partition coefficient to form regression equations, and inclusion of an electronic parameter such as Hammett sigma or quantum mechanically derived electronic charges on the chelating atoms did not improve the correlation coefficient significantly. The method was extended to model the separation efficiencies of mercaptobenzothiazoles (MBT) and aminothiophenols (ATP) used in the flotation of lead and zinc ores, respectively. Five principal components were found to explain 99% of the data variability in each series. A three-parameter equation with correlation coefficient of 0.985 and a two-parameter equation with correlation coefficient of 0.926 were obtained for MBT and ATP, respectively. The amenability of separation efficiencies of chelating collectors to QSAR modeling using PCs based on topological indices might lead to the selection of collectors for synthesis and testing from a virtual database.

INFLUENCE OF PULP PULSATION ON THE BATCH FLOTATION OF GALENA

An experimental investigation into the effect of pulp pulsation on the flotation of an artificial mixture of galena and quartz has been conducted. Laboratory experiments have shown that the recovery and kinetics of lead flotation can be improved significantly by pulsating the pulp, while the grade and selectivity of the lead remained virtually unchanged. Pulsation of the pulp led to structural change of bubbles. An increase in the number of bubbles and a decrease in the bubble size resulted in an increase in the gas holdup in the pulp and a likely increase in the bubble-particle collision rate. Likewise, the viscosity of the slurry could be decreased by means of pulp vibration, enhancing mass transfer between the slurry phase and froth phase. At the same time, the energy input by pulp pulsation could help to lower the energy barrier for bubble-particle attachment by lowering the induction time of the attachment. In addition, non-selectively attached particles could be stripped off the bubble surface by the inertial forces acting on the particles during pulsation of the pulp. Image analysis of the froth showed that pulsation resulted in a more stable froth, with more numerous, smaller bubbles with higher loadings.