Direct Leaching Research Articles

Design Criteria for the Construction of Energy Storage Salt Cavern Considering Economic Benefits and Resource Utilization

Underground salt caverns have been widely used for oil and gas storage and have attracted increasing attention. The construction design of salt caverns is directly related to the final storage capacity, economic benefits, and resource utilization. However, due to the numerous combinations of multi-stage process parameters involved in the construction design, it is difficult to optimize them individually through indoor experiments and numerical simulations. In this regard, this paper attempts to put forward the basic principles of cavern construction design criteria with economic benefits and resource utilization as indicators. Firstly, 1258 groups of cavern construction process parameters were randomly generated under certain basic rules, including inner tube depth, outer tube depth, oil pad depth, duration, and water injection flow rate, for five direct leaching stages. Then, the cavern capacity, economic benefit, and rock salt resource utilization corresponding to these process parameters were obtained through batch processing using single-well salt cavern leaching simulation software (SSCLS). Finally, the influence laws of the distance between the inner tube and oil pad and lifting heights, and the rates of the inner tube and oil pad on the expected economic benefits and salt resource utilization, are discussed. In the actual project, it is recommended to increase the distance between the inner tube and the oil pad, increase the ratio of oil pad lifting height to duration, and use the appropriate lifting height to obtain greater expected revenue and resource utilization. This work will improve the efficiency and scientificity of cavern construction design, which is of great significance in guiding the construction and design for energy storage in salt caverns.

Evaluation and Current State of Primary and Secondary Zinc Production—A Review

This article presents the history of zinc, its production and demand. The quantity of zinc production, both primary zinc from ores and concentrates, and secondary zinc from scrap and zinc-rich waste, was discussed. A comprehensive economic analysis covers zinc prices in the years 1960–2021. The basic methods of obtaining zinc from ores, including pyrometallurgical (Imperial Smelting Process ISP, Kivcet, Ausmelt) and hydrometallurgical (roasting–leaching–electrowinning RLE, atmospheric direct leaching ADL, Engitec Zinc Extraction EZINEX, zinc pressure leach) and their short characteristics, are presented. The global zinc market and the main areas of its application were analyzed. Technologies used for the recovery of zinc from scrap are discussed along with their characteristics. Galvanized steel is the main source of secondary zinc, both in the galvanizing process and in the remelting of galvanized steel. It can be easily recycled with other scrap steel in the electric arc furnace (EAF) for steel production. Currently, with high volatility in the price of zinc, as well as its natural resources in the earth’s crust, recycling is an important activity, despite the fact that zinc concentrates have a relatively constant chemical composition, while the resulting zinc waste contains zinc in varying amounts.

Effect of Hydrophobicity of Ionic Liquids on the Leaching Selectivity of Platinum from a Spent Automotive Catalyst

Recycling platinum group metals from secondary resources such as spent automotive exhaust catalysts is promising for the circular economy. However, the selective separation of Pt over massive amounts of impurity metals such as Mg and Al is particularly challenging. In this study, non-aqueous direct leaching of platinum from a spent automotive catalyst (SAC) using hydrophobic ionic liquids, namely, trihexyl(tetradecyl)phosphonium chloride (P66614Cl) and trioctyl(dodecyl)phosphonium chloride (P88812Cl) with the aid of pre-loading hydrochloric acid/hydrogen peroxide was proposed. The more hydrophobic P88812Cl exhibited more efficient and selective leaching of Pt over Mg and Al. The recovery of Pt from the metal-loaded P88812Cl, and the reusability of the IL for SAC leaching were also demonstrated.

Iron Extraction from South African Ilmenite Concentrate Leaching by Hydrochloric Acid (HCl) in the Presence of Reductant (Metallic Fe) and Additive (MgSO4)

The high content of iron in ilmenite ore poses a great challenge, particularly in the synthesis of titanium-containing products due to high susceptibility of iron (Fe) to corrosion. Direct leaching of ilmenite ore in hydrochloric acid (HCl) encouraging Fe dissolution was investigated. The influence of variable parameters, the use of additives, and the addition of metallic iron powder were studied to establish the optimum leaching parameters. The results showed that ilmenite with the particle size distribution of +150 µm yielded better efficiencies when leaching was performed with an acid concentration of 7.5 M and a solid-to-acid ratio of 1:10 at 90 °C. An agitation speed of 450 rpm yielded a superior Fe extraction of about 92.32% and a 2.40% titanium (Ti) loss. The addition of both metallic Fe and the MgSO4 additive significantly enhanced Fe dissolution and decreased Ti recovery in a leach solution. It was found that leaching under optimum conditions produced a solid residue with 1.37% Fe impurity while 98.63% was extracted. The leached residue was comprised of 91.4% TiO2 rutile phase and contained a high content of the ilmenite FeTiO3 (4.37%) and SiO2 (2.23%) impurities, while Al2O3, MgO, MnO2, CaO, V2O5, MnO2, and Cr2O3 were below 0.13%. The high TiO2 content in the leached residue makes it suitable for use as feed in the production of synthetic rutile. The leaching kinetics of Fe dissolution was found to conform to the shrinking core model, where diffusion through the product layer is rate controlling. The calculated activation energy according to the Arrhenius equation was 19.13 kJ/mol.

Study on modes of occurrence and enhanced leaching of critical metals (lithium, niobium, and rare earth elements) in coal gangue

This paper first studied the occurrence modes of critical metals (lithium: Li, niobium: Nb, and rare earth elements: REY) in coal gangue (CG) by sequential chemical extraction, and the distribution and enrichment degree of critical metals in different density and size fractions were also studied, and then CG was pretreated with intercalation and calcination to strengthen the extraction of critical metals. The results of sequential chemical extraction showed that the residual fraction was the dominant occurrence mode of critical metals. The enrichment ratios of different density fractions were 0.57–1.23 for Li, 0.03–1.34 for Nb, 0.69–1.36 for REY, and different size fractions were 0.89–1.23 for Li, 0.46–2.21 for Nb, 0.63–1.17 for REY. Density separation and particle size classification can be used to enrich critical metals. Direct leaching of 3 M hydrochloric acid resulted in 5.76% Li and 0.74% Nb recovery. Further, the leaching efficiency of Li was improved to 17.83% using pretreatment of intercalation, and the crystal structure of kaolinite was retained in the leaching residue. In comparison, calcination pretreatment led to 92.74% Li recovery, while the leaching efficiencies of Al and Fe were correspondingly improved. This paper can provide a theoretical foundation and technological support for the extraction of critical metals from CG.

Efficient Recovery of the Combined Copper Resources from Copper Oxide Bearing Limonite Ore by Magnetic Separation and Leaching Technology

The reserve of the copper-oxide-bearing limonite ore (COBL ore) in Yulong Copper Co., Ltd. is up to 20 million tons with 1.79% of copper content. The characters of the copper resources in the COBL ore are high-proportioned oxidation state (99.98%) and combined state (84.83%). The combined copper oxide is mainly copper-oxide-bearing limonite, which has a copper content of more than 78%. Because of the high altitude and average annual temperature of 15 °C in Tibet, fire leaching cannot be adopted. The leaching efficiency of copper from COBL ore using direct leaching of sulfuric acid is only 40%, which is greatly influenced by temperature and time. Based on the characteristics of COBL ore, a novel combined method of magnetic separation and individual leaching has been proposed to efficient recover copper resources. Experimental results show that the magnetic concentrates and tailings were obtained by magnetic separation of COBL ore at 0.6 T with the yields were 59.65% and 40.35%, respectively. Due to the obvious leaching properties difference of the magnetic concentrates and tailings, individual leaching process routes were used to treat them. The magnetic concentrate was leached with stirring for 3 days at room temperature (20 °C), and the magnetic tailing was easily leached for 4 h at 40 °C. The recovery efficiency of total copper was 72%, which was about 32% higher than that of the single leaching of the COBL ore. The method proposed in this study achieves environmentally friendly, low energy consumption, and efficient extraction of refractory copper oxide ore.

Effect of surface modification of polyester fabric on microfiber shedding from household laundry

PurposeSynthetic textile materials are noted as one of the major contributors to microfiber pollution through laundry. Though many research works evaluated microfiber pollution, the solutions provided to control microfiber shedding are meager. The existing products collect or filter the microfiber from laundry effluent and restrict the direct leaching. However, no methods were proposed to effectively reduce the shedding from the textile itself.Design/methodology/approachThis research is aimed to analyze the influence of surface modification of polyester knitted textiles by sodium hydroxide, on microfiber shedding. Response surface methodology was adapted to optimize different treatment parameters (alkali concentration, treatment time and temperature).FindingsThe results show that the sodium hydroxide concentration and treatment time had a negative correlation with microfiber shedding reduction. Whereas, treatment temperature had a positive correlation with microfiber shedding reduction. The statistical analysis revealed that 0.4 M concentration, 90°C temperature and 24 min of treatment time was the best process condition for minimum microfiber release. The same was confirmed with a practical experiment and a significant reduction of 80.63% in microfiber shedding after alkali treatment was found.Originality/valueAlkali treatment of different knitted polyester fabrics with various knit structures and mass per square meter showed a significant reduction in microfiber shedding. The repeated laundry performed for 20 washes with surface-modified samples showed a significant reduction in microfiber release at every wash cycle and ensured the longevity of the effect.

Economic analysis of rare earth element processing methods for Mountain Pass ore

The increasing demand for rare earth elements (REE) in highly strategic sectors such as hybrid electric vehicles, renewable energy, communication technologies, and the defense industry has made REE production methods important. In recent years, studies on REE have increased rapidly; however there are many studies on REE, there are not enough studies in the literature on the economics of the beneficiation process. RE ores are generally leached after beneficiation by gravity, magnetic and flotation methods. In this study, direct leaching, attrition scrubbing and leaching, and flotation and leaching methods are evaluated economically based on the data for the Mountain Pass (MP) facility. According to the sensitivity analysis performed as a function of capital cost (CAPEX), operating cost (OPEX), and rare earths oxides (REO) recovery, REO recovery has the biggest effect on the process economies. Although the flotation method entails additional investment and operating costs, transportation and operating costs are significantly reduced result of the higher grade and recovery. The flotation method is more profitable, with larger NPV and IRR values and a shorter payback period.

Comparative Studies on Hydrometallurgy and Pyrometallurgy to the Extraction of Beryllium from Decrepitated Beryl from Wadi El Gemal Region, Eastern Desert, Egypt

Comparative Studies on hydrometallurgy and pyrometallurgy to the extraction of beryllium from decrepitated beryl from Wadi El Gemal Region, Eastern Desert, Egypt are carried out in this work in short, useful summary. Beryl is one of the important ore mineral constituents of the mineralized zones at Wadi El Gemal area, Eastern Desert, Egypt. Extraction of beryllium from beryl ore involves both (a) pyrometallurgical techniques through roasting of beryl with Na2SiF6 followed by leaching of roasting beryl with commercial (dilute) H2SO4 (b) hydrometallurgical processing techniques of beryl through direct leaching with H2SO4 to produce directly sulfate liquor. 88% of beryllium was extracted by roasted beryl with Na2SiF6 at 850°C for 1.5 hours and leaching the roasted product with 25% H2SO4 AT 85 ° C for 3/4 hour. On the other hand, 95 % of beryllium was extracted directly after 20 minutes by direct leaching at 280°c with 10 mol/liter H2SO4. Therefore, this work sheds light on the study of the factors affecting both forms of beryllium extraction.

Leached nitrate under fertilised loamy soil originates mainly from mineralisation of soil organic N

Animal manures are suspected to be a major source of nitrate leaching due to their low nitrogen use efficiency (NUE) by crops. However, actual measurements of nitrate leaching from animal manure under field conditions are scarce. In an on-farm field trial in Switzerland over 2.5 years, we used 15N labelling to trace the fate of N from cattle slurry in the soil-plant system and to test whether more nitrate was leached from slurry than from mineral fertiliser. The experiment was conducted on two neighbouring fields with loamy soil in an agricultural area of the Swiss midlands, where nitrate levels in the groundwater are persistently high. Both fields followed the same crop rotation (silage maize – winter wheat – grass-clover), but shifted by one year. We compared three fertiliser treatments: Control (Con), 15N mineral fertiliser (Min), and 15N cattle slurry (Slu). In order to provide a comprehensive fertiliser N balance over several years, we traced the labelled fertilisers into crop biomass, soil, and leached nitrate. In the year of application, 15N recovery in crops was 45–47% for Min, but only 19–23% for Slu. Complementary to this finding, recoveries in soil were greater for Slu than for Min, despite greater NH3 emissions from Slu. Fertiliser recovery in the succeeding crops was small (< 4.6% of the originally applied fertiliser N in the first residual year and < 2.4% in the second) and similar for both fertilisers. Depth translocation of fertiliser N was marginal, with the majority of 15N in soil still in the top 0.3 m after 2.5 years. Along with higher recoveries in soil for Slu, we found significantly more slurry N than mineral fertiliser N lost through leaching. However, less than 5% of cumulated amounts of nitrate leaching over the three crops, which reached up to 205 kg nitrate-N ha-1, originated from direct leaching of the labelled fertilisers. Our findings suggest that most nitrate leaching originated from the mineralisation of soil N.

Leaching of Gold and Silver from a Complex Sulfide Concentrate in Copper-Tartrate-Thiosulfate Solutions

The leaching behaviors of gold and silver from a complex sulfide concentrate in copper-tartrate-thiosulfate solutions were investigated in this paper. Experimental parameters, including temperature, initial pulp pH, and concentration of copper, tartrate, and thiosulfate, were systematically studied. The copper-tartrate-thiosulfate leaching system exhibits promising performance in dealing with the complex sulfide concentrate. Thiosulfate consumption could be greatly reduced due to the in-situ generation of thiosulfate derived from sulfur or disulfide ions. Increasing the temperature and concentrations of copper, tartrate, and thiosulfate can promote gold leaching. A low tartrate content and a moderate increase in temperature and copper and thiosulfate concentrations can accelerate silver leaching, but a higher temperature or copper content may depress silver extraction. A suitable pulp pH is critical for gold and silver leaching in copper-tartrate-thiosulfate solutions. An extraction of 74.50% Au and 36.33% Ag was obtained with the direct leaching of the concentrate, while the percentages can be significantly increased up to 82.60% and 70.38%, respectively, when leaching the calcine following the oxidative roasting pretreatment. The recycling of leaching solutions demonstrates that a suitable free tartrate content is of great significance in maintaining the stability of the novel system. As an alternative to the traditional cyanide or copper-ammonia-thiosulfate leaching processes, the copper-tartrate-thiosulfate leaching system provides an environmentally friendly, nontoxic, and relatively low-cost method for gold and silver leaching from ores or concentrates.

Leaching of yttrium and europium from fluorescent lamp phosphor powder using nitric acid: kinetics and optimization

ABSTRACT Rare earth elements (REEs) are recognized as one of the most critical elements and considered essential for the development of sustainable energy technologies, lasers, electronics, alloys, catalysts, and other applications. In order to meet the increasing demand to ensure those essential usages, the supply of REEs is however limited from the primary sources. Recovery of REEs from secondary resources, such as discarded electrical and electronic equipment, referred to as electronic waste (e-waste) is consequently crucial. Fluorescent lamps among such wastes are of special concern due to the high quantity of REEs present in them. This article discusses a potential hydrometallurgical method for recovering REEs (yttrium and europium) from fluorescent lamp waste via nitric acid leaching. The effects of leaching factors including acid concentration and temperature were examined. To assist the leaching yields of those metals, the effect of alkaline fusion was specifically examined. The kinetics of both processes, direct- and alkaline fusion-assisted leaching were investigated based on the shrinking core model. In the case of direct leaching, the metal dissolution was controlled by surface chemical reaction with apparent activation energies of 43.9 and 48.1 kJ/mol for yttrium and europium, respectively, in the temperature range 303–333 K. Pretreatment, viz., the alkaline fusion changed the subsequent leaching mechanism to diffusion control with apparent activation energies of 7.4 and 9.5 kJ/mol for the respective metal in the same temperature range. Optimization of the alkaline fusion-assisted acid leaching process was undertaken by applying response surface methodology (RSM) based on the central composite design (CCD). Under the alkaline fusion conditions such as soda/fluorescent powder mass ratio of 0.65 g/g, temperature of 1223 K and reaction time of 7200 s followed by acid leaching (nitric acid concentration – 3.5 M, temperature-323 K, pulp density – 30 g/L and leaching duration – 1200 s), the leaching efficiencies of 92.6% and 99.5% for yttrium and europium, respectively, could be achieved.

Extraction of Rare-Earth Elements from Silicate-Based Ore through Hydrometallurgical Route

The European Union and several countries/regions classified the rare-earth elements (REEs), such as lanthanum, cerium, neodymium, and scandium, as critical due to the risk of supply interruption. For this reason, the growing demand for REEs has resulted in forgotten reserves receiving economic interest. So, the search for new sources and the development of chemical process is important, such as silicate-based ore. Since there is almost no literature on the extraction of REEs from this source, a new approach was developed in the present study. Direct leaching and acid baking were studied using sulfuric acid. The effect of the acid concentration, temperature, solid-liquid ratio, oxidizing/reducing medium, and acid dosage were studied. Results showed that the extraction of REEs achieved up to 80% at 90 °C in oxidizing medium, and scandium and iron achieved 13.5% and 65.0%, respectively. For the acid baking experiments, the results were better than direct leaching for REEs at over 85%. The scandium leaching rate was lower than direct leaching. On the other hand, the extraction of iron was lower in acid baking than direct leaching. The iron and scandium extraction rates were higher in lower temperatures (&lt;200 °C) and acid dosages, achieving 50% and 6.3%, respectively. Future studies should explore thermal treatment before acid leaching.

Recovery Cu ad Ce from copper slag by using flotation and chemical methods

Slag, which is now considered as a secondary source, attracts attention due to the metallic values they contain. Copper slags having economical operating grades in terms of Zn, Cu, Ce, Ni, Co, Mo, and V show complex mineralogy due to its thermal processing history. Various methods for the recovery of copper and other precious metals from this byproduct have been studied, and there are many operating flotation plants in industrial scale. However, in cases where physical or physo-chemical mineral processing methods are not sufficient, hydrometallurgical methods are applied. In this study, flotation was applied to copper slag with 0.9% Cu grade. Copper recovery and concentrate grade were found as around 40 and 22.7% Cu, respectively. Further, metal extractions were examined by conducting direct leaching tests. Gel formation was observed, when H2SO4was used in the leach process, therefore, the leach parameters were investigated using HCl. Cu and Ce extractions of around 70-80% were reached as a result of a test performed for 2 hours with the addition of 25 g/L H2O2 at a rate of 20-25% solids and 300 g/L HCl concentration. When the relationship of dissolved metals with each other wasexamined, dissolution behavior of Cu and Ce was quite similar to Al dissolution (R2=0.9).

Selective separation of base metals and high-efficiency enrichment of precious metals from scrap copper anode slime

In this study, a clean and effective integrated process for selectively separating base metals and high-efficiency enriching precious metals from scrap copper anode slime was investigated. O2-enriched roasting combined with H2SO4-NaCl leaching was used in the decopperized process to avoid the scattered distributions of As & Sn and enhance the selective leaching efficiencies of Cu, Ni, and As. The obtained leachate containing Cu, Ni, and As can be sent to the existing electrolyte purification process. Then, alkaline xylitol leaching was applied to selectively separate Pb greatly avoiding its negative impact on the subsequent process, and Pb-containing leaching solution was used to prepare lead electrodeposition electrolyte. Sn was recovered in the form of sodium stannate by soda roasting-alkaline leaching. Results of scale-up experiment showed that the direct leaching efficiencies of Cu, Ni, As, Pb, and Sn under the optimum conditions were 99.3 %, 96.9 %, 98.6 %, 98.1 %, and 98.9 %, respectively. Ag and Au were maximally remained in the residue. Furthermore, three leaching solutions generated in the proposed process can be treated in an environmentally friendly way. Therefore, it was a high-efficiency and eco-friendly method for SCAS to realize the comprehensive utilization of the base metals and reduced risks of secondary pollution.

Promoted acid leaching of Zn from hazardous zinc-containing metallurgical dusts: Focusing on transformation of Zn phases in selective reduction roasting

As hazardous wastes from metallurgical processes, high zinc dust (HZD) was considered as a potential alternative resource of Zn. The acid leaching of HZD was a principal step and therefore was highly relevant to efficient utilization. Unfortunately, zinc ferrite (ZnFe2O4) in HZD was difficult to be leached by hydrometallurgical methods. Therefore, in this study, to enhance the recovery of zinc, selective reduction roasting was used to decompose ZnFe2O4. The transformation of Zn phases during the reduction process was investigated by thermodynamic calculation, XRD analysis. The results showed that ZnFe2O4 was converted into acid-soluble ZnO in H2 at 600 °C, while the further reduction of ZnO to Zn(g) and the volatilization of ZnO was avoided. However, the soluble ZnSO4 in HZD was also reduced to acid-insoluble ZnS. Consequently, a novel method of reduction roasting combined with water leaching was proposed to pretreat HZD. In water leaching, the solid-liquid ratio was1:10 (g/mL). When H2SO4 concentration was 80 g/L, the leaching ratio of Zn was 94.24% after pretreatment, higher than that of direct leaching of HZD (87.76%). Therefore, acid leaching of Zn was promoted by water leaching and reduction roasting pretreatment, which will bring considerable benefits to the high zinc dust recycling enterprises.

Recovery of rare earth elements from coal fly ash using enrichment by sodium hydroxide leaching and dissolution by hydrochloric acid

ABSTRACT The most common method to recover valuable rare earth elements from the potential resource coal fly ash is direct leaching in acidic solution. However, the extraction of rare earth elements using only acid leaching is not sufficiently effective due to the presence of Al-Si-based glasses. In the present work, pre-treatment using NaOH leaching was investigated to reduce the content of Al-Si-based phase before subjecting to acid leaching. The treatment at conditions NaOH 5 M, at 100 °C and for 120 min could dissolve 53% of the ash sample, and the rare earth contents were enriched from 267 mg/kg to 556 mg/kg. The NaOH leaching promote the further extraction of rare earth elements in all of acid medias (HCl, HNO3 and H2SO4), and HCl was the most effective leachant. More than 90% of rare earth elements was dissolved in 5 M HCl solution at 50°C after 120 min. The combination of NaOH treatment and HCl leaching presented an effective approach to extract rare earth elements from coal fly ash with enrichment of rare earth element contents, increase the leaching efficiency and avoidance of harmful impacts to the environment due to the conventional pre-treatment using alkaline roasting.

Experimental and statistical modelling on copper recovery from low-grade mixed sulphide-oxide ore by direct leaching

ABSTRACT This study aims to investigate the recovery of copper from a low-grade source of an ore deposit in sulphuric acid solution. In this study, a full-factorial statistical design model and Artificial Neural Networks (ANN) are applied to develop an analytical approach to understand the dissolution parameters better. Temperature, time, particle size, and acid concentration were selected as the leaching parameters. The experimental design set by the full-factorial model was again trained with the ANN model to compare the predicted results. The correlation coefficients (R2) in both the models are 0.9962 and 0.99815, respectively, indicating both the models offered a well-aligned explanation for the variability of the experimental values to the predicted values. Therefore, the outcome of this study demonstrates that the proposed design model results are well-aligned with the experimental results, and 88% of copper metal can be recovered at a condition of 120 minutes, 90°C temperature, the particle size of -75 + 63 µm, and acid concentration of 1M.

Hydrochloric Acid Leaching of Philippine Coal Fly Ash: Investigation and Optimisation of Leaching Parameters by Response Surface Methodology (RSM)

Coal Fly Ash (CFA) is a hazardous waste from coal-fired power plants, but has increasingly become a popular supplementary cementitious material for cement in the construction industry. As a secondary resource of REE, its main advantage lies in its fine particle size that eliminates the need for costly and energy-intensive comminution. In this study, the potential of CFA from the Philippines as a secondary REE resource was investigated by direct leaching of REEs with hydrochloric acid (HCl). The CFA sample came from a coal-fired power plant with a circulating fluidized bed combustion (CFBC) technology. For the leaching tests, the effects of HCl concentration, leaching time and leaching temperature on REE extractions were elucidate optimized via Response Surface Methodology (RSM). The RSM results showed that the optimum leaching occurred at 3M HCl, 65 °C and 270 min with extractions of Nd, Er, Eu, Tb and Dy at 70.8%, 76.34%, 88.02%, 90.01% and 73.38%, respectively. According to these results, the CFA from the Philippines is a promising secondary resource of REE and the extraction method employed was effective in achieving a relatively high REE dissolution. Moreover, the empirical model that was established accurately predicted the dissolution of REE with an accuracy of 98.20%, 96.66%, 97.09%, 98.17% and 97.78% for Nd, Er, Eu, Tb and Dy, respectively.

Efficient Recovery of Vanadium from High-Chromium Vanadium Slag with Calcium-Roasting Acidic Leaching

High-chromium vanadium slag (HCVS) is an important by-product generated during the smelting process of high-chromium-vanadium-titanium-magnetite. Direct acid leaching and calcium-roasting acid leaching technology were applied to recover vanadium and chromium from HCVS. The effects of experimental parameters on the leaching process, including concentration of H2SO4, reaction temperature, reaction time, and liquid-to-solid ratio, were investigated. The XRD and UV-Vis DRS results showed that vanadium and chromium existed in low valence with a spinel structure in the HCVS. The Cr-spinel was too stable to leach out; no more than 8% of the chromium could be leached out both in the direct acid leaching process and calcium-roasting acid-leaching process. Most low valence vanadium could be oxidized to high valence with calcium-roasting technology, and the leaching efficiency could be increased from 33.89% to 89.12% at the selected reaction conditions: concentration of H2SO4 at 40 vt.%, reaction temperature of 90 °C, reaction time of 3 h, liquid-to-solid ratio of 4:1 mL/g, and stirring rate of 500 rpm. The kinetics analysis indicated that the leaching behavior of vanadium followed the shrinking core model well, and the leaching process was controlled by the surface chemical reaction, with an Ea of 58.95 kJ/mol and 62.98 kJ/mol for direct acid leaching and roasting acid leaching, respectively.