Study on Recovering Zinc from Gossan

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.

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.

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.

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.

Two-stage leaching of zinc and copper from arsenic-rich copper smelting hazardous dusts after alkali leaching of arsenic

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.

An analysis of the composition of brass metallurgical waste was carried out using X-ray fluorescence analysis (XRF). The results showed that the copper content in the slag reaches 15 wt. %, and the zinc content – 83 wt. %. Sulfuric acid leaching was performed to separate zinc from the metallurgical dust. Optimal process parameters were selected: leaching duration – 60 minutes, sulfuric acid concentration – 0.1 M. After sulfuric acid leaching, the acid solution was subjected to electrochemical treatment to recover copper and zinc, and the copper cake (copper in unoxidized form) was subjected to copper-ammonia leaching for 40 minutes. The concentration of copper in the copper-ammonia solution reached 35 g/L. At the final stage, solvent extraction of copper from the copper-ammonia leach solutions was carried out, for which the most effective extractant was selected. Extractants of different nature and classes were studied: D2EHPA (a strong acidic organophosphorus extractant), DХ510А and LIX54 (belonging to the class of β-diketones). The concentration of extractants ranged from 50 to 100%, the diluent was kerosene. Copper stripping from the copper-ammonia extract was performed using 2M sulfuric acid. The best extractant was found to be LIX54 at 50% concentration in kerosene. The final stage was copper electrowinning from the stripping solutions at a current density of 3 A/dm, with a current efficiency of 65%. Based on the conducted research, a process flowsheet was developed for the recovery of copper and zinc from brass metallurgical dust.

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.

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 comprehensive recycling of aluminum electrolysis cell waste barrier material is urgent. This study focuses on the sulfuric acid leaching of waste barrier material, systematically examining the effects of factors such as reaction temperature, liquid-to-solid ratio, sulfuric acid concentration, and reaction time on the leaching of elements like lithium, aluminum, sodium, and silicon. The experimental results show that under the conditions of 0.9 mol/L sulfuric acid concentration, a liquid-to-solid ratio of 20:1, a reaction temperature of 90 °C, and a reaction time of 1.5 h, the leaching rates were 84.5% for lithium, 85.6% for aluminum, 98.5% for sodium, and 4.8% for silicon. The sulfuric acid leaching process of the waste barrier material follows a shrinking core model and is controlled by internal diffusion. The apparent activation energies for the leaching reactions of lithium, aluminum, and sodium were 4.29 kJ/mol, 8.99 kJ/mol, and 9.11 kJ/mol, respectively. The selective leaching of lithium, sodium, and aluminum from silicon was successfully achieved in the sulfuric acid leaching of the waste barrier material.

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.

In recent years, recovering zinc from low grade oxidized zinc ore has been a matter of discussion. In this paper, the investigation for recovering zinc and iron from gossan ores was carried out. In generally, the conventional physical separation methods, such as flotation, gravity separation and magnetic separation, were used to treat the low grade ore, and flotation is the most commonly used method for beneficiation and pretreatment of oxidized zinc minerals. However, gravity separation and magnetic separation are used to deal with the gossan ores because of the special zinc oxide ore bearing iron 40.2% in this study. The effects of grinding fineness and magnetic intensity were investigated, respectively. The results show that the grade of zinc and iron increase slightly, and the best result is that the grade of zinc and iron are 13.35% and 40.25%, the recovery of zinc and iron are 53.45% and 52.19%.

<p class="1Body">The sulfuric acid leaching of titanium from titanium-bearing electric furnace slag (TEFS) was investigated under different experimental conditions. In the sulfuric acid leaching process, the M<sub>x</sub>Ti<sub>3-x</sub>O<sub>5</sub>(0≤x≤2) and diopside could react with sulfuric acid. The optimum conditions of sulfuric acid leaching process were particle size at < 0.045mm, sulfuric acid concentration at 90 wt.%, acid/slag mass ratio at 1.6:1, feeding temperature at 120 °C, reaction temperature at 220 °C, reaction time at 120minute, curing at 200°C for 120 minute. The [TiO<sub>2</sub>] concentration of the water leaching was 150 g/L, and leaching temperature at 60℃for 120 minute. Ti extraction could reach 84.29 %. F of titanium-bearing solution was 2.15, and the Ti<sup>3+</sup>/TiO<sub>2</sub> of the titanium-bearing solution was 0.068. The TiO<sub>2</sub> content of the leaching residue was 18.32 wt.%. The main mineral phases of the leaching residue were calcium sulphate, spinel, diopside and little M<sub>x</sub>Ti<sub>3-x</sub>O<sub>5</sub>.</p>

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

The article describes results of research on copper recovery from low-grade copper ore by heap bioleaching method. The objects of investigation are ores of the Benkala deposit. Results of chemical analysis by atomic-emission spectrometric method and chemical phase analysis present ores element composition and identify copper and iron forms existing in the ores. Ores were subjected to bioleaching by chemolithotrophic bacteria, which oxidizes sulfur and iron compounds. Acidithiobacillus ferrooxidans FT-24 and BF, Acidithiobacillus thiooxidans BS, Acidithiobacillus ferrivorans SU-8 and Sulfobacillus thermosulfidooxidans ST-12 strains were used in the research. It was modeling the process of bacterial heap leaching of low-grade ore in percolation columns. Comparison of efficiency of sulfuric acid and bacterial leaching in percolation columns shows advantage of bioleaching. The copper yield was 47 % at using conventional sulfuric acid leaching, while utilizing bacterial leaching increased the copper recovery up to 86 % during 90 days of the experiment. The values of oxidation-reduction potential (ORP) at sulfuric acid and bacterial leaching of copper ore presented are in accordance with yield of copper. The values of ORP at standard sulfuric acid leaching are between 330-360 mV, at bacterial leaching more high and are 480-550 mV due to high content of ferric iron. During the extraction of copper, the effect of the organic reagent LIX 984N in concentrations 50 and 250 mg/L on the activity of microorganisms was studied. As a result, the extraction reagent has a little effect on the growth of microorganisms and the extraction of copper. Without adding the extraction reagent, the total copper recovery was about 83 %, while the addition of the extraction reagent with concentration 250 mg/L decreased it to 81 %. So using bioleaching technology allows deep processing of copper ore due to oxidation of copper sulfide minerals.