Zn Grade Research Articles

Beneficiation of Oxidized Lead-Zinc Ores by Flotation Using Different Chemicals and Test Conditions

This study presents the recovery of oxide-carbonate lead-zinc ores using different chemical reagents under different test conditions by flotation. The run-of-mine oxide-carbonate Pb-Zn ore contains 9.05% Pb and 11.97% Zn with a major mineralization of smithsonite and cerussite. First experimental work was grinding tests to reduce -106 μm size fractions with two 15-minute grinding stages. Pre-flotation tests gave similar results for Pb, but neither pH, ZnSO4, CuSO4 nor collectors like AERO 3477, 3501 and 8651 made favorable effect on the recovery of Zn. Therefore, the studies were emphasized for recovering Pb concentrate by changing the amount of chemicals used, such as CMC, Na2S, AERO promotor. After six stage of flotation with 350 g/t KAX, 275 g/t AERO 407, 7500 g/t Na2S, 1000 g/t CMC, the concentration of 70.93% Pb with 71.56% recovery was achieved. The 91.51% Zinc remained in the tailing with 14.66% Zn grade.

Simulation-aided studies of heavy-media separation in Angouran lead and zinc ore

ABSTRACTPre-concentration of lead and zinc ore in Calcimin (public Co) which is done using heavy-medium cyclones (HMC) undergoes lots of inefficiencies and not in the desired manner. Considering the fact that feed material is variable in the case of Zn–Pb grades, it is important to make HMC plant flexible when facing different ore types. The present study aims to prepare a predicting tool which can determine the operational parameters of HMC plant such as density separation and predict the properties of the final product e.g. Zn–Pb grades and recoveries using existing models. After determining the feed ore characterizations including particle size distribution, XRF, XRD, and high-density sink-float tests, an existing empirical model for high-density HMC was used. Inserting the data obtained from ore characterization tests in existing models and solving necessary calculations makes it possible to predict the behavior of HMC plant at different ore types and select the most favorable operational conditions without doing a mass of experimental work. An optimal flowsheet of HMC plant was proposed as an example for the studied sample which shows the possibility of producing a pre-concentrate with Zn and Pb grades of 36.60 and 3.24%, respectively, with the recoveries of 93.66 and 90.31%.

Cr(III) and Fe(II) recovery from the polymetallic leach solution of electroplating sludge by Cr(III)-Fe(III) coprecipitation on maghemite

The separation of Cr(III) from the polymetallic leaching solution is a typical problem encountered during the recycling of electroplating sludge. In this paper, a novel method was proposed to recover Cr(III) and Fe(II) synchronously by forming the Cr(III)-Fe(III) coprecipitates on the surface of maghemite (γ-Fe2O3) fine particles with Fe2(SO4)3 solution as the supplemental Fe source. The results showed that high Fe(III) to Cr(III) ratio in solution promoted the hydrolysis and precipitation of Cr(III). In addition, the maghemite particles, served as the crystal nuclei, could induce the formation of the core-shell structured Cr(III)-Fe(III) coprecipitates on its surface and accelerate the sedimentation of the coprecipitates in the magnetic field. By using a two-stage Cr(III)-Fe(III) coprecipitation process, the recoveries of Cr and Fe reached 96.17% and 99.39%, and the grades of Ni, Cu, and Zn in the coprecipitates were at 0.41%, 0.38%, and 0.22%, respectively. The obtained coprecipitates can be recycled as the feed material of chromium smelting after heat treatment. This method is simple and efficient for high-concentration Cr3+ solution treatment, which is beneficial for the sustainable development of resources and environment.

A comparative study of two-stage flotation of Zn and Pb oxide minerals using anionic, cationic, and mixed (cationic/anionic) collectors

A lead-zinc carbonate ore sample containing 2.5% Pb and 9.39% Zn was used in this research work. The sample was prepared from the Darreh-Zanjir mine located in the Yazd province (Iran). Influences of the influential factors on flotation of smithsonite and cerussite were investigated. Among the different parameters involved, dosages of the dispersant, depressants, sulfidizing agent, and collectors de-sliming prior to lead or zinc flotation were essential for the effective recovery and grade of the Zn and Pb flotation concentrates. In addition, the anionic, cationic, and mixed (cationic/anionic) collectors were employed for flotation of smithsonite. The results of reverse and cumulative flotation of both Zn and Pb were relatively low in comparison with the direct process without depressant. Flotation of smithsonite using mixed collectors (Armac C+KAX) showed promising results. Also dosages of chemicals in the cleaning stage for the Zn and Pb concentrates were optimized, and finally, after the cleaner stage for both lead and zinc, a cerussite concentrate with Pb grade and recovery of 49.82% and 60.06%, respectively, and smithsonite concentrate with Zn grade and recovery of 35.47% and 68.56%, respectively, were obtained under the optimal conditions. Furthermore, kinetics of Zn and Pb oxide mineral flotations in the rougher and cleaner stages were studied, which showed that some kinetics models, especially the classical first-order model, could predict the flotation behaviour of the Zn and Pb oxide minerals.

Sodium carbonate effects on the flotation separation of smithsonite from quartz using N,N′-dilauroyl ethylenediamine dipropionate as a collector

To improve the flotation separation efficiency of smithsonite from quartz using N,N′-dilauroyl ethylenediamine dipropionate (DNET) as a collector, sodium carbonate was used as a regulator. Single-mineral flotation results indicated that more than 97.36% of smithsonite was floated when DNET was used as a collector at a concentration of 200 mg/L. However, artificially mixed minerals were poorly separated during the activation of quartz with Zn(II) due to smithsonite dissolution. Sodium carbonate could inhibit the dissolution of smithsonite and thus improve the flotation separation of smithsonite and quartz. The results of artificially mixed mineral separation showed that a concentrate with 89.7% Zn recovery and 36.6% Zn grade could be obtained when the DNET concentration was 50 mg/L at a slurry pH of 10.45. To clarify the action of sodium carbonate during the separation process, zeta potential measurements, Fourier transform infrared (FTIR) spectroscopic analysis, and zinc ion concentration measurements were carried out. In addition, the results indicated that sodium carbonate could inhibit the solubility of smithsonite, and thus, the activation of quartz was prevented during the interaction of CO32− with dissolved Zn(II), thereby improving the separation results.

Use of ZnSO4 and SDD mixture as sphalerite depressant in copper flotation

The depression performance and mechanism of the zinc sulfate (ZnSO4) and sodium dimethyl dithiocarbamate (SDD) combined depressant with butyl xanthate (BX) as collector in a mixed Cu-Zn sulfide floation system were fully investigated using micro-flotation experiments, zeta-potential determination, and local electrochemical impedance spectroscopy (LEIS) analysis. The micro-flotation experimental results indicated that the combined depressant has a significant selective depression effect on sphalerite but has slight effect on chalcopyrite. The depression performance of the combined depressant strictly depends on the dosage of ZnSO4 + SDD and flotation pH. However, excessive SDD is detrimental to chalcopyrite flotation. The optimum conditions for flotation separation of chalcopyrite from the Cu-Zn sulfide were as follows: combined depressant dosage of 10−4 mol/L (ZnSO4: SDD = 3:1), BX dosage of 10−5 mol/L, and flotation pH of 10. Under such conditions, a copper concentrate assaying 30.21% Cu with 86.79% recovery was achieved from the Cu-Zn sulfide flotation system, and this concentrate had a very low Zn grade of 4.20% with 5.48% recovery. Both the zeta-potential and LEIS results confirmed that the presence of the combined depressant can significantly prevent the adsorption of BX onto the sphalerite surface, especially for an alkaline pH range of 9–12; however, the adsorption of BX onto the chalcopyrite surface is not affected. The LEIS provides a powerful tool for observing the adsorption of flotation reagents on mineral surface.

Study on the pre-treatment of oxidized zinc ore prior to flotation

The pre-treatment of zinc oxide bearing ores with high slime content is important to ensure that resources are utilized optimally. This paper reports an improved process using hydrocyclone de-sliming, dispersion reagents, and magnetic removal of iron minerals for the pre-treatment of zinc oxide ore with a high slime and iron content, and the benefits compared to traditional technologies are shown. In addition, this paper investigates the damage related to fine slime and iron during zinc oxide flotation, the necessity of using hydrocyclone de-sliming together with dispersion reagents to alleviate the influence of slime, and interactions among hydrocyclone de-sliming, reagent dispersion, and magnetic iron removal. Results show that under optimized operating conditions the entire beneficiation technology results in a flotation concentrate with a Zn grade of 34.66% and a recovery of 73.41%.

Lead and zinc geochemical behavior based on geological characteristics in Parkam Porphyry Copper System, Kerman, Iran

Parkam (Sarah) porphyry system is located on the metallogenic belt of Kerman, Iran. Due to existence of some copper-rich resources in this region, finding out the exact statistical characteristics such as distribution of data population, mean, variance and data population behavior of elements like Cu, Mo, Pb and Zn is necessary for interpreting their geological behavior. For this reason, precise calculation of statistical characteristics of Pb and Zn grade datasets was performed and results were interpreted geologically. The natures of Pb and Zn distributions were initially identified and their distributions were normalized through statistical treatment. Subsequently, the variograms were calculated for each exploration borehole and show that both Pb and Zn geochemical variates are spatially correlated. According to the similarity of the behavior of Pb and Zn in these calculations, it is decided to measure their exact behavior applying K-means clustering method. K-means clustering results show that the Zn grade varies linearly relative to that of Pb values and their behavior is similar. Based on the geochemical behavior similarity of Pb and Zn, throughout the pervasive secondary hydrothermal activity, they are remobilized in the similar manner, from the deep to the shallow levels of the mineralization zones. However, statistical analysis suggests that hydrothermal activity associated with secondary waters in Parkam is effective in remobilizing and enriching both Pb and Zn since they have similar geochemical characteristics. However, the process does not result in generation of economic concentrations.

Mineralogy and mineral chemistry of the metamorphosed and precious metal-bearing Ming deposit, Canada

The Ming deposit in the northwestern Canadian Appalachians is a metamorphosed, bimodal-mafic, precious-metal (Au, Ag)-bearing and Cu-rich volcanogenic massive sulfide (VMS) deposit consisting of several, spatially proximal lenses in a sericite- to chlorite-altered rhyodacitic footwall. The various lenses (1807 Zone, 1806 Zone, Ming South Up Plunge, Ming South Down Plunge, and the Lower Footwall Zone) have variations in Cu, Au, Ag, and Zn grades that reflect varying physico-chemical conditions of ore formation. This study describes the complex ore mineralogy of the orebodies and constrains the genesis of the deposit using field methods, mineralogy, whole rock sulfide geochemistry, and micro-analytical methods.The orebodies are predominantly pyrite, chalcopyrite, with lesser sphalerite, pyrrhotite, and trace galena and arsenopyrite, with the exception of the Lower Footwall Zone, which consists of a high temperature (>320°C) chalcopyrite–pyrrhotite–pyrite±cubanite assemblage. The other orebodies (1806, 1807, Ming South Up Plunge and Down Plunge) contain trace amounts of tellurides (hessite, altaite, tsumoite, unnamed bismuthtelluride), sulfosalts (Ag-poor and Ag-rich tennantite–tetrahedrite, meneghinite, AgSb phases, stannite), and precious metal phases (electrum, AgHg±Au alloys). The 1807 Zone is enriched in Te, Bi, and Se, whereas the 1806 Zone is telluride-free and contains As, Sb, Hg, Au, and Ag. Mineral chemistry of sphalerite shows strong variations in Fe content (1.12–11.04wt.%). Intermediate Fe (4.33–6.33wt.% Fe) and Fe-rich (7.327–11.04wt.% Fe) sphalerite are common in all orebodies, whereas Fe-poor sphalerite (1.12–3.57wt.% Fe) occurs exclusively in the 1807 and 1806 zones. Tennantite-tetrahedrite is typically Ag-poor (0.25–2.19wt.%) in the 1807 Zone, but is significantly enriched in Ag (up to 29.3wt.%) in the 1806 Zone. Galena in the 1807 Zone and Ming South orebodies is commonly myrmekitically intergrown with tellurides and has high concentrations of Te, Bi, Se, and Ag. In contrast, galena in the 1806 Zone is less enriched in Te, Bi, and Se, but high in Ag.Variations in mineralogy, epithermal elements (As, Bi, Hg, Sb, Se, Sn, Te) and precious metal (Ag, Au) content, and mineral chemistry between the different orebodies indicate that they were formed from predominantly reduced, acidic hydrothermal fluids with varying ƒTe2/ƒS2, ƒSe2/ƒS2, and mBi/mSb ratios as temperatures steadily decreased from >300°C to <260°C during ore formation. In the 1807 and Ming South orebodies, late-stage deposition of Te-, Bi-, Se-, Ag-rich galena and tellurides occurred prior to precipitation of Ag-poor tennantite-tetrahedrite, whereas in the 1806 Zone hydrothermal fluids low in ƒTe2/ƒS2, ƒSe2/ƒS2, and mBi/mSb favored the precipitation of Te-, Bi-, Se-poor galena and Ag-rich tennantite-tetrahedrite. A magmatic source for epithermal elements and precious metals is suggested and was part of the depositional history of the Ming deposit.

Separation and recovery of ZnS from sulfidized neutralization sludge via the hydration conversion of CaSO4 into bulk CaSO4·2H2O crystals

Hydrothermal sulfidation–flotation is considered to be a feasible means for the heavy metal recovery from neutralization sludge (NS) generated by the lime-neutralization process in our previous studies. However, the as-generated sulfides prefer to aggregate on the surface of CaSO4 during hydrothermal sulfidation, which causes difficulties in the sulfides recovery via flotation. In this study, the hydration of CaSO4 was proposed to promote the separation of sulfides (mainly ZnS) in sulfidized NS. When hydration was performed on pure CaSO4, the rod-shaped CaSO4 with an average length of 10 μm was converted into CaSO4·2H2O with an increased average length of 100–400 μm. Based on this, hydration was introduced into the sulfidized NS, and the CaSO4 in the sulfidized NS was converted into bulk CaSO4·2H2O crystals. Sieving experiments toward the hydration product show that a total of 90.66% of the Zn was concentrated in a −48 μm sieving product. To further improve the zinc grade of the −48 μm sieving product, a flotation test was conducted, during which a concentrate with Zn grade of 37.11% was obtained and Zn recovery reached 66.87%. In contrast, when flotation was directly performed on sulfidized NS, the Zn grade in concentrates and Zn recovery were only 21.96% and 37.43%, respectively. Therefore, the hydration of CaSO4 in sulfidized NS can effectively promote the separation and recovery of ZnS.

Investigation of Cu, Mo, Pb, and Zn geochemical behavior and geological interpretations for Parkam porphyry copper system, Kerman, Iran

A well-known algorithm of clustering is K-means by which the data are divided into K classes based upon a distance criterion. In this study, we apply the K-means method for classifying data derived from exploration boreholes in the Parkam deposit. The optimum K has been calculated and then the data have been clustered and the relative geochemical behavioral characteristics analyzed. The criterion used for determining the optimum K ranged in number of classes from K = 3 to K = 10, and afterwards, we analyzed the derived classifications in order to choose the optimum K. Results showed that class numbers of K = 3 in the case of Cu and Mo, K = 4 in the case of Cu and Pb, and K = 3 in the case of Cu and Zn were optimized class numbers. After clustering, the increasing Cu grade values resulted in a significant increase in Mo grades, a significant decrease in Pb grades followed by an increase, and the Zn grades varying comparable to Pb. With regard to the relationships between these elements, it can be concluded that the meteoric waters promoted the mobilization of Pb and Zn from the potassic zone to the phyllic, but the meteoric waters were not effective enough to cause the mobilization of Cu, and this element together with Mo remained immobile.

Comprehensive recovery of metals from cyanidation tailing

Cyanidation tailing is the residue produced in gold plants which use cyanidation to extract gold. It can be used as a secondary resource to recover residual metals that are of great economic value. The cyanidation tailing investigated in this paper was obtained from Shandong Province, China. It contained valuable metals such as chalcopyrite, galena, sphalerite and pyrite. In this study, alkaline sodium hypochlorite was used as a regulator in the pretreatment stage. It was proved that the sodium hypochlorite played two roles in the flotation pulp: oxidant and pH regulator. On one hand, sodium hypochlorite oxidized cyanide to cyanate, eliminating the negative effect of residual cyanide towards the environment. On the other hand, with the pH of flotation pulp exceeding 10, sphalerite and pyrite were depressed enormously, which was beneficial to the recovery of chalcopyrite and galena. With the Cu–Pb bulk flotation flowsheet, the cyanidation tailing was processed to obtain qualified Cu concentrate with grade of 13.17% and recovery of 70.00% compared with the original Cu grade of 0.21%. The Cu–Pb tailing was processed to obtain qualified Zn concentrate with grade of 34.72% and recovery of 69.58% compared with the original Zn grade of 0.33%, constituting the comprehensive recovery routing for the cyanidation tailing.

Smithsonite Flotation from Zinc Oxide Ore using Alkyl Amine Acetate Collectors

Flotation of smithsonite from tailing coming from a cerussite flotation circuit was investigated using Armac C and Armac T as alkyl amine acetate collectors. The results showed that the dosage of Na2S, collector type and dosage, collector's mixture ratio, and desliming are the main affecting parameters on the Zn grade and recovery of smithsonite flotation concentrate. The contact angle measurement on smithsonite and calcite minerals indicated that Armac T acts more selectively than Armac C and enhances the Zn grade of the concentrate. At the mixture ratio of Armac C:Armac T, 1:6, the collectors' consumption are decreased significantly and an optimal concentrate with 94.4% recovery and 42.3% Zn grade is obtained without any gangue depressant reagents. Nevertheless, using NaOH as a pH regulator, the dosage of Na2S is decreased significantly but the flotation results are affected negatively. The recovery of Zn in the concentrate is more sensitive to the slime than Zn grade. The Zn minerals lost in the tailing are mostly in the form of very fine disseminated particles. About 11% of smithsonite with particle sizes finer than 10 µm is lost in the slime. Overall Zn flotation recovery is estimated to be about 84%.

Flotation Experimental Study of High-Sulfur Cu-Zn Ore in Yunnan

The experimental ore belongs to high-sulfur Cu-Zn poly-metallic ore, the mineral composition of ore is complex, cupriferous pyrite most are disseminated. The Cu grade of raw ore is 0.76%, Zn grade is 1.48% and S about 25.02%. on the basis of best reagent system and control flotation time, the author did closed circuit process. the results of closed circuit process as follows: the copper concentrate grade is 20.35% and recovery rate is 86.081%; the zinc concentrate grade is 54.32% and recovery rate is 84.20%.

Structural and Thermoelectric Properties of Zn-Doped Cuprous Aluminate Delafossite

The polycrystalline Zn-doped CuAlO2 ceramics were obtained by solid state reaction method. The mixture of high purity grade of CuO, Al2O3 and Zn (CH3COO)22H2O powders was ground and then pressed by using uniaxial pressure. The obtained pellet was sintered in air at 1,150 °C for 24 h. XRD patterns showed the crystal structure of the as-sintered CuAl1-xZnxO2 (0x0.05) belonging to rhombohedral, space groupalong with the CuO and CuAl2O4 phases. From SEM micrographs, the grain size slightly decreases when the amount of Zn in CuAl1-xZnxO2 samples increases. Hall mobility and hole concentration of the as-sintered samples were obtained from Hall effect measurements in van der Pauw configuration at room temperature. The electrical resistivity measurements as a function of temperature were performed. The variation of Seebeck coefficient and power factor as a function of temperature was also investigated. In the present study, the Zn-doped CuAlO2 ceramics probably have advantages for use in high temperature thermoelectric devices.

Study on Pre-Concentration and Calcium Removal of Low Grade Zinc Oxide Ore

The zinc oxide sample from the Lanping mine located in the west of Yunnan province，China, is characterized with low grade, high calcium content, high oxidation rate and heavy slime. According to the properties of the ore, a new dressing-metallurgy combined method is proposed to process the ore and the beneficiation part is discussed by this paper. Based on the regulation of reagent scheme, flotation flowsheet of pre-desliming and two rougher is conducted to obtain a zinc concentrate that Zn grade is 18.72% at Zn recovery of 90.88%. The calcium removal rate of zinc concentrate reaches to as high as 78.15% for the flotation operation, which could lay the foundation for the next acid leaching process.

Experimental Research of Mineral Processing on a Certain Lead-Zinc Oxidize Ore

A certain multi-metallic lead-zinc oxide ore contains 1.09%Pb，8.39%Zn． The oxide rate of lead and zinc is 96.34% and 98.15% respectively．In order to recover valuable minerals from the oxide ore comprehensively, the authors adopted priority flotation processing in this experimental research. At last, the lead concentrate with Pb grade of 27.41% and the zinc concentrate with Zn grade of 24.14 % were respectively achieved. The overall recovery of lead was 93.66%，and the overall recovery of zinc was 90.74%.

Flotation of zinc oxide ore using cationic and cationic-anionic mixed collectors

Flotation of smithsonite coming from a cerussite flotation circuit was investigated using Armac C and Armac T as cationic collectors and their mixtures with potassium amyl xanthate (KAX) as an anionic collector. Under optimum condition using Armac T with sodium hexamethaphosphate as a de- pressant, a smithsonite concentrate with 42% Zn and 89.6% recovery is obtained while in the presence of sodium silicate depressant Armac C produces a concentrate containing 40.8% Zn with 92.5% recovery. However, Armac T acts a little more selectively than Armac C but their mixture with KAX act inversely. The KAX-Armac C mixed collector improves the Zn grade and recovery of the flotation concentrate while KAX-Armac T does not. At the mixture ratio of KAX - Armac C 2:1, the amine collector consump- tion is decreased and an optimal concentrate with 94% recovery and 43% Zn grade is obtained without any gangue depressant reagents.

Impact of Pre-Desliming on Flotation of Low-Grade Zinc Oxide Ore from Lanping

Zinc oxide ore from Lanping contained large amount of fine slime which badly deteriorated flotation performance, and it complexly intergrowed with plenty of alkaline gangues with calcium and magnesium. In order to investigate the impact of pre-desliming on flotation, experiments of direct flotation and pre-desliming flotation were carried out in the work. Finally, a zinc recovery of 88.73% with a Zn grade of 13.93% was obtained in direct flotation test, and a higher zinc recovery of 95.25% with a Zn grade of 13.56% was achieved by pre-desliming flotation. Pre-desliming flotation test was superior to direct flotation test, because of an increase of nearly 7% in zinc recovery under the condition of similar Zn grade. At the same time, dosage of sulfurization agent used in pre-desliming flotation was lower by 46.67% than that in direct flotation, which obviously improved the availability of sulfurization agent. So both product index and economic benefit were enhanced after pre-desliming.

An Integrated Process for Innovative Extraction of Metals from Kupferschiefer Mine Dumps, Germany

AbstractDumps in the Central German Mansfeld mining district have been investigated by a multi‐disciplinary research team to develop a holistic concept for the recovery of base and trace metals from low‐grade Kupferschiefer ore. Geochemical and mineralogical data from representative sampling indicate that the low‐grade Kupferschiefer dumps contain significant amounts of Cu (0.6 %) and Ag (up to 800 ppm), besides exploitable grades of Pb and Zn. Additionally to flotation and hydrometallurgy, first shake flask bioleaching experiments on otherwise untreated Kupferschiefer ore yield Cu recovery rates of up to 95 %. A total of 900 t Cu and 5 t Ag contained in black copper can thus be produced from the pilot dump of 100 000 m3 low‐grade Kupferschiefer ore. Additionally, large volumes of limestone waste, separated from the dumps by sensor‐based sorting, can be utilized as aggregate in the construction industry.