EXTRACTION AND SEPARATION OF COPPER AND ZINC FROM METALLURGICAL DUSTS AND SLAGS OF BRASS PRODUCTION BY ELECTROCHEMICAL AND EXTRACTION METHODS

Gossan discarded from mining processes result in metal resource wastage, and its long-term stacking causes environmental hazards. Therefore, this article considers zinc-containing gossan as the research object. The ore was roasted to prepare primary zinc ferrite products and sulfuric acid leaching was performed for purification. Then, XRD analysis was performed to characterize the purified products. The results indicated that the effect of sulfuric acid concentration on the purification of the products was related to its zinc ferrite content. Furthermore, the effect of leaching temperature on the purification of zinc ferrite products was related to sulfuric acid concentration; the lower the sulfuric acid concentration, the more considerable the effect of leaching temperature. The conditions suitable for purifying the products through sulfuric acid leaching are as follows: sulfuric acid concentration of 140 g/L, liquid–solid ratio of 4:1, leaching temperature of 80 °C, leaching time of 120 min, and stirring speed of 300 rpm. This article determines the factors affecting the purification of zinc ferrite by sulfuric acid leaching along with the optimal purification conditions. The findings presented herein provide a theoretical foundation for the development of new processes for preparing zinc ferrite, which has considerable industrial application value.

The main purpose of these studies was to study the chemical and mineralogical composition of the ore of the Kenshoky-1 deposit with subsequent search experiments on the extraction of copper from ore material by sulfuric acid leaching. During the research it was established that the ore belongs to the type of oxidized cuprous sandstone and along with copper it contains a significant amount of lead and a small amount of zinc. The main minerals of copper and lead in this ore are carbonate – malachite and cerussite. A fairly significant amount of the lead phosphate mineral pyromorphite is present. The main minerals in the waste rock are quartz, orthoclase, and clinochlorine. Iron, according to diffractometric analysis, is bound in clinochlore. The search experiments on copper leaching from the mentioned ore by sulfuric acid solutions showed that even at room temperature and sulfuric acid concentration in the solution of 11.3 g/dm3 extraction of copper in aqueous solution of more than 60% is achieved, and the main technological factors influencing the process of copper leaching are duration, concentration of sulfuric acid in the initial solution, intensity of agitation and the ratio of solid and liquid phases. The concentration of copper in the resulting pregnant solution was, depending on the concentration of sulfuric acid from 6.32 to 10.15 g/dm3 , which is insufficient for further processing of solution by electrolysis, therefore, productive solutions must be additionally saturated with copper by returning pregnant solution to the head of the process. Lead behavior during prospecting studies was not studied, but due to low solubility of lead sulfate in aqueous solutions, the main part of lead should remain in the cake

In recent years, recovering zinc from zinc calcine with high iron has been a matter of discussion. In this paper, sulfuric acid leaching was carried out to assess the effect of several parameters on zinc and iron extraction in zinc calcine with high iron in which the grade of zinc and iron is 53.90% and 19.38%, respectively. Parameters, such as stirring speed, sulfuric acid concentration, liquid to solid ratio and leaching time, were investigated. The results show that leaching time has done nothing to the leaching rate, but has great influence on leaching efficiency. Liquid to solid ratio and sulfuric acid concentration have significant influence on leaching results, and stirring rate has not obvious influence on leaching results. Under the condition of 120g/L sulfuric acid, 6:1 liquid to solid ratio, 55°C leaching temperature and 120min leaching time, the recovery of zinc and iron is 82.24% and 9.64%, respectively. It is obvious that ZnO in zinc calcine is easy to dissolve in acidity solution, which shown in two aspects: high leaching rate and high leaching speed. ZnO can be dissolved entirely in sufficient sulfuric acid in 10min.

Natural graphite is classified as a strategic critical raw material by the European Union due to its strategic significance. The demand is steadily increasing, but the current purification methods are not environmentally sustainable, highlighting the need for greener alternatives. This study investigates the effects – sulfuric acid (H2SO4) leaching, thermal treatment, and their combination – on natural graphite from a flotation pilot plant. In the sulfuric acid leaching, the influences of reaction time (60–300 min), temperature (70–100°C), sulfuric acid concentration (0.5–3 mol/L), and liquid-to-solid ratio (10–20 mL/g) were systematically studied using experimental design. Leaching effectively removed 91.4% iron and 52.9% aluminum, but was ineffective against silicon-containing phases. A short 15 min thermal treatment at 2400°C in an argon atmosphere using induction annealing eliminated most silicon phases and other impurities, although residual 2.65 mg/g iron and 1.27 mg/g silicon remained. The combined approach reduced iron and silicon contents to 0.24 and 0.25 mg/g, respectively, and increased the carbon content from 78.4 to 97.6 wt% to near commercial battery-grade levels (98.0 wt%). Additionally, the combination-treated graphite exhibited the lowest degree of structural defects, offering a more sustainable route for purifying natural graphite to high-purity levels compared to conventional halogen-containing techniques.

Rare earth element recovery in molten salt electrolysis is approximately between 91 and 93%, whereof 8% is lost in waste molten salt slag. Presently, minimal research has been conducted on the technology for recycling waste rare earth molten salt slag, which is either discarded as industrial garbage or mixed with waste slag into qualified molten salt. The development of a new approach toward the effective treatment of rare earth fluoride molten salt electrolytic slag, which can recycle the remaining rare earth and improve the utilization rate, is essential. Herein, weak magnetic iron separation, sulfuric acid leaching transformation, water leaching, hydrogen fluoride water absorption, and cycle precipitation of rare earth are used to recover rare earth from their fluoride molten salt electrolytic slag, wherein the thermodynamic and kinetic processes of sulfuric acid leaching transformation are emphatically studied. Thermodynamic results show that temperature has a great influence on sulfuric acid leaching. With rising temperature, the equilibrium constant of the reaction gradually increases, and the stable interval of NdF3 decreases, while that of Nd3+ increases, indicating that high temperature is conducive to the sulfuric acid leaching process, whereof the kinetic results reveal that the activation energy E of Nd transformation is 41.57 kJ/mol, which indicates that the sulfuric acid leaching process is controlled by interfacial chemical reaction. According to the Nd transformation rate equation in the sulfuric acid leaching process of rare earth fluoride molten salt electrolytic slag under different particle size conditions, it is determinable that with the decrease of particle size, the reaction rate increases accordingly, while strengthening the leaching kinetic process. According to the equation of Nd transformation rate in the sulfuric acid leaching process under different sulfuric acid concentration conditions, the reaction series of sulfuric acid concentration K = 6.4, which is greater than 1, indicating that increasing sulfuric acid concentration can change the kinetic-control region and strengthen the kinetic process.

The statistical modeling of the equilibria in the solvent extraction of copper from sulfuric acid and ammoniacal sulfate solutions by mixtures of LIX 64N and SME 529 has been carried out. This study employed an incomplete three-level factorial design involving four variables,viz., the copper and sulfuric acid or ammonia concentrations in the aqueous phase and the concentrations of the two extractants in the organic phase. The equilibrium models showed that the active oximes in LIX 64N and SME 529 are completely compatible as regards the extraction of copper without any synergistic effects between the reagents being observed. The percentage of copper extracted increases with increased extractant concentration in the organic phase, but is affected adversely by increased sulfuric acid and copper sulfate concentrations in the aqueous phase. As expected, the equilibrium extraction of copper is influenced by the diluent, with higher copper extractions being attained with mixtures of LIX 64N and SME 529 in MSB-210 as compared with Escaid 100, a diluent having a higher aromatic content and giving rise to a lower degree of oxime aggregation. The copper loading efficiency of the reagents is high for extractions from ammoniacal solution — 86.6 pct and 89.4 pct copper in the diluents MSB-210 and Escaid 100, respectively, at the center points of the experimental designvs 47.0 pct and 43.3 pct in the case of sulfuric acid solution extractions. Copper loading efficiencies greater than 100 pct are attainable in ammoniacal media, particularly in situations where the aqueous copper concentration is greatly in excess of the combined extradant concentration. This overloading is attributable to the uptake of ammonia by the hydroxyoximes and can be expected to increase with increasing aromatic content of the diluent, as in the case of Escaid 100, as a result of decreased oxime aggregation.

The content of zinc in copper smelter slags obtained from pyrometallurgical copper production is comparable to the content of this metal in zinc ores. Therefore, these slags are considered a valuable secondary resource for zinc recovery. At the same time, the features of the mineralogical composition of the slag make the extraction of zinc from it very problematic. Most of the zinc is concentrated in the refractory zinc ferrite (ZnFe2O4). To avoid the formation of a viscous pulp when leaching copper smelter slag with an aqueous solution of sulfuric acid, in this work, the slag was leached with sulfuric acid also in isopropanol and n-pentanol, under the following conditions: 0.5 M H2SO4, pulp density 50 g/L, magnetic stirrer rotation speed 600 rpm. The influence of the duration and temperature of leaching milled (≤100 μm) copper smelter slag of the Balkhash copper smelter on the extraction of zinc into solution was investigated. It was found that the maximum zinc recovery into an aqueous solution was 75 ± 2% at 363 K and 210 min. Replacing water with isopropanol or n-pentanol led to an increase in zinc recovery to 82 ± 2% at 210 min and a lower temperature (353 K) than in an aqueous environment. An increase in temperature to 383 K during leaching in n-pentanol made it possible to extract 92 ± 2% of zinc. A shrinking core model was used to describe the kinetics of the zinc leaching process. It was found that the limiting stage of the process under all investigated conditions is the chemical leaching reaction. Some kinetic characteristics of the leaching process were calculated, in particular, the apparent reaction rate constants, as well as the activation energy.

An experimental study of the decomposition process of Kovdorsky apatite with certain particle sizes, in a 1 dm3 batch reactor with stirring of the reaction mixture, initial phosphoric acid concentration 17% by weight, in the system: Apatite-H3PO4 – H2SO4-H2O. Sulfuric acid was introduced in stoichiometric amount at the beginning of the process. The process was carried out at a ratio L:S 2.5:1, at a temperature of 78– 82 °C. The process was monitored by the method of joint determination of sulfuric and phosphoric acids by titrimetric analysis of the composition of the reaction mixture in the presence of methyl orange, and then phenolphthalein. Determined the concentration of phosphoric acid, the concentration of excess or deficiency of sulfuric acid and the concentration of monocalcium phosphate gel were determined. The acidity of the reaction mixture was recorded according to the indications of a Ph-meter (pH-105 MA with a combined glass electrode – ESK-10603). In the system: Apatite-H3PO4-H2SO4-H2O, when analyzing the obtained experimental values of these parameters, it was shown that during the process the pH of the reaction mixture rises to pH 6.3 for about 30 minutes, and then decreases to pH 4.5–5. Comparison of the dependence of pH values on the concentration of sulfuric acid in an aqueous solution of phosphoric acid and the pH values of the reaction mixture with the corresponding concentrations of sulfuric and phosphoric acids shows that the pH value of the reaction mixture is determined by the presence in the mixture between weft product - monocalcium phosphate gel. The process of decomposition of apatite in the system Apatite-H3PO4-H2SO4-H2O proceeds in several stages. At the beginning, a rapid decomposition of phosphate raw materials with phosphoric acid and a corresponding change in the pH of the reaction mixture, associated with the accumulation of monocalcium phosphate in solution, takes place. Then there is a slow decomposition of monocalcium phosphate with sulfuric acid and a corresponding decrease in pH to 4.5–5. The pH value at the end of the process is determined by the physicochemical properties of phosphoric acid present in the reaction mixture. At all stages of the process there is a constant increase in the concentration of phosphoric acid associated with the decomposition of monocalcium phosphate gel.

Shaimerden oxidized zinc ore containing more than 21% wt.% zinc is studied in this work. When the first research on hydrometallurgical processing of this ore was conducted the authors of this research determined the following parameters for ore leaching: the size of ore reduction, sulfuric acid concentration, temperature, and leaching duration. These tests serve as a basis for zinc recovery from the given ore with a size of – 2.5 mm and a regime is recommended providing zinc recovery into solution at the level 93.5%. However, the production regime proposed for sulfuric-acid leaching of this zinc ore requires further development as this process is characterized by inadequate production, high energy consumption for heating leaching pulp, and considerable sulfuric acid consumption. In order to avoid these disadvantages the authors develop a production regime of four-stage sulfuric-acid leaching of Shaimerden ore – 1.0 mm in size. The regime developed in this article provides zinc recovery from ore into solution at the level 94.65% compared with the regime of sulfuric-acid leaching developed previously. At the same time, consumption of sulfuric acid for leaching is reduced by 22%, the leaching period is six times shorter, and energy consumed for heating leaching pulp is lower by approximately a factor of three.

In this paper, zinc ferrite was prepared from zinc calcine by sulfuric acid leaching, and its properties were studied by the methods of XRD, XRF and SEM/EDS. The results show that sulfuric acid leaching of zinc ferrite from zinc calcine is mainly affected by the initial concentration of sulfuric acid and leaching temperature. If the initial concentration of sulfuric acid and leaching temperature are too low, the soluble oxides in zinc calcine can not be completely dissolved. The prepared zinc ferrite was not suitable for application for too many impurities. However, the higher initial concentration of sulfuric acid and leaching temperature are, the lower recovery of zinc ferrite is. The suitable conditions for preparation of zinc ferrite from zinc calcine by sulfuric acid leaching are liquid-solid ratio 7:1, stirring speed 400rpm, sulfuric acid concentration 160g/L, leaching temperature 85°C, leaching time 120min. The particle size of zinc ferrite is fine, most of them is between 2~5μm. The impurity content of zinc ferrite is less. Zinc ferrite particles are composed of spherical and irregular shape, and they are agglomerate and encase each other.

The present study was designed to investigate the effect of mercuric chloride administration on copper, zinc, and iron concentrations in the liver, kidney, lung, heart, spleen, and muscle of rats. The results showed that after dose and time exposure to mercuric chloride, the concentration of mercury in the six tissues was significantly elevated. Data showed that there were no interaction between mercury and tissue iron. There was a considerable elevation of the content of copper in the kidney and liver. The most significant changes in the copper concentration took place in the kidneys. About a twofold increase in the copper content of the kidney was noted after exposure to mercuric chloride (3 mg and 5 mg/kg). Only slight elevations in the copper content occurred in the liver especially in high dose and longer exposure time. In the remaining organs, the copper content was not changed significantly (p > 0.05). The most significant changes in the zinc concentration took place in liver, kidney, lung and heart (5 mg/kg). Marked changes in kidney zinc concentrations were observed at any of the specified doses. Zinc concentrations were significantly increased in kidney of rats sacrificed 9-48 h after s.c. injection of HgCl2 (5 mg/kg); in liver obtained from rats at 18, 24 or 48 h after injection; and in lung after 24 or 48 h of treatment. The heart and spleen zinc concentrations were elevated at 24 and 48 h after injection of HgCl2 (5 mg/kg), respectively. The results of this study implicate that effects on copper and zinc concentrations of the target tissues of mercury may play an important role in the pathogenesis of acute mercuric chloride intoxication.

Large amounts of manganese-rich slag were generated from smelting low-grade manganese ore containing high iron and high phosphorus in electric furnace or blast furnace, which contain 25–40 mass% Mn. This paper describes an efficient hydrometallurgical process for the utilization of high manganese slag instead of conventional usage of pyrometallurgical route for manganese recovery. Pressure acid leaching of manganese-rich slag using air as pressurized gas is studied systematically. The effects of mass ratio of manganese-rich slag to pyrite, concentration of sulfuric acid, leaching temperature, partial air pressure, leaching time, and liquid-solid (L/S) ratio on the separation of manganese with other main impurity components were investigated comprehensively. Under the following optimal process conditions, weight of manganese-rich slag powder,100 g; mass ratio of manganese-rich slag to pyrite, 20:1; sulfuric acid concentration of 140 g/L; air pressure of 0.6 MPa; leaching temperature of 140 °C; leaching time of 1 h; L/S of 7:1; ore size of 75–80 μm; and stirring rate of 600 rpm, the extraction rate of Mn reached 94%, and the dissolution percentage of Si, Al, and Fe are as low as 2%, 6%, and 10%, respectively. The leaching slurry presented a good solid-liquid separation characteristic. The silicic acid formed during sulfuric acid leaching manganese silicates was converted to filterable dehydrated SiO2 particles which precipitated in the residue under the condition of higher temperature with higher pressure. The effect of adding pyrite may be to reduced iron ions and decreased the output of iron hydroxide colloid and also reduced MnO2 to manganese sulfate in sulfuric acid leaching. Thus the formation possibility of colloid such as silica gel and iron hydroxide was decreased, and higher yield of manganese sulfate could be achieved.

The production of ferroalloys occupies a leading position in the economy of the Republic of Kazakhstan. With the depletion of rich in the content of the initial manganese raw materials, a problem arose that it was necessary to involve technogenic products, i.e. poor sub-standard ores and wastes produced by ferroalloys. From the point of view of environmental protection of regions producing manganese alloys, the current tasks are: the introduction of new efficient dust collection processes; issues of cleaning of waste gases and sludges; neutralization and utilization of waste. The creation of rational technological schemes for the utilization of dispersed materials that contain the target element of the ferroalloy to be melted is an economically viable and environmentally sound measure for increasing the profitability of production. Sulfuric acid leaching is the main operation of most schemes of hydrometallurgical processing of manganese-containing raw materials. Hydrogen peroxide, metallic iron, pyrite concentrate, sulfur dioxide, sulfite-bisulphite solutions used as the reducing agent. The article presents the results of hydrometallurgical processing of manganese dusts produced by the silicomanganese of the Aksu Ferroalloy Plant. To convert the solution of psilomelane (MnOMnO2) – an oxide compound, in which manganese is represented in silicomanganese dust, the presence of a reducing agent in the sulfuric acid solution, which is used as a pyrite concentrate, is necessary. The effect of pyrite on the reduction of manganese dioxide during leaching with sulfuric acid studied by mathematical planning of the experiment by a probabilistic-deterministic method. The determining factors of the leaching process are: temperature, duration of the process, amount of pyrite added, concentration of sulfuric acid. On the basis of significant equations of partial dependence, a mathematical model for the leaching of manganese dust by sulfuric acid in the presence of pyrite is compiled in the form of a generalized equation: Yрасч = 3,7•10-6(0,9399х1+5,1847)(-13,761х22+62,507х2+23,402)(-0,7429х32+14,143х3+23,4)(-00071х42+1,466х4+18,323). On the basis of the obtained equation, the optimal conditions for manganese leaching into a sulfuric acid solution are selected: temperature 700 °C, duration 3 hours, sulfuric acid concentration 5 %, additive of pyritic concentrate 90 % of the dust weight. The degree of manganese extraction was 95.8 %.

In this paper, the conventional physical separation method such as flotation, gravity separation, magnetic separation, alkaline leaching and sulfuric acid leaching were studied. The effects of grinding fineness, amount of agent, magnetic intensity, roasting temperature, roasting time, the leaching agent and leaching time on the leaching of zinc were investigated, respectively. The results show that the leaching rate of zinc is below 50% in the conventional alkaline leaching, and the leaching rate of zinc is below 85% and the leaching rate of iron is above 35% in sulfuric acid leaching. Compared with XRD pattern of the raw ore, the different diffraction peaks of smithsonite is off in alkaline leaching products. In sulfuric acid leaching, the different diffraction peaks of smithsonite are off in the leaching products when sulfuric acid concentration is less than 60 g/L. After 60 g/L, the different diffraction peaks of smithsonite and siderite are off in the leaching products.

In this paper, sulfuric acid leaching was carried out to assess the effect of several parameters on metal extraction in a low grade complex gossan ore in which the grade of zinc and iron is 13% and 40.2%, respectively. Parameters, such as sulfuric acid concentration, liquid to solid ratio and leaching temperature, were studied. The results show that the zinc leaching rate is almost 80%, while the iron leaching rate is about 45% used strong acid with 200g/L. It can be seen from the results that sulfuric acid leaching could not effectively recover zinc from gossan ores studied in this paper because of iron dissolving greatly.