Roasting-water Leaching Research Articles

A Comparison Study on the Recovery of REEs from Red Mud by Sulfation Roasting–Water Leaching and Citric Acid Leaching

In this study, the recovery of rare earth elements (REEs) from red mud (bauxite residue) was explored through a combination of citric acid leaching and sulfation roasting–water leaching processes, introducing an innovative approach to the field. The research uniquely investigates the influence of citric acid on the leaching behavior of REEs and impurities in both untreated red mud and red mud subjected to sulfation roasting, providing a direct comparison of these methodologies. A novel aspect of this study is the evaluation of solvent extraction efficiency using DEHPA, highlighting the selective recovery of REEs over impurities from both citric acid and water-leaching solutions. Furthermore, a comprehensive phase analysis using X-ray diffraction (XRD) was conducted to track the transformations of minerals during the sulfation roasting process, an original contribution to the literature. The findings revealed that over 85% of REEs and major elements such as Fe, Al, Ca, and Ti dissolved in water after sulfation at 105 °C, while iron and titanium dissolution significantly decreased following roasting at 725 °C. Importantly, terbium, neodymium, and gadolinium extraction efficiencies were notably affected by roasting temperature. Citric acid leaching results demonstrated that the direct leaching of red mud leads to higher leaching efficiency than leaching it after the roasting process. Solvent extraction demonstrated lower terbium and neodymium recovery from citric acid solutions compared to water leaching solution. Finally, stripping experiments illustrated that 6M H2SO4 solution is capable of stripping more than 80% of rare earth elements, except terbium.

Selective lithium recovery from waste lithium-ion batteries by H2SO4 roasting focused on process intensification and conversion mechanism

The purpose of this study is to prove the feasibility, efficiency and selectivity of lithium recovery from waste lithium-ion batteries (LIB) via sulfuric acid roasting-water leaching by thermodynamic analysis and experimental verification. The feasibility was proved by Gibbs free energy calculations and experiments. Then, the effects of H2SO4 concentration, roasting temperature and roasting time were assessed on the leaching rate of valuable metals. The results indicate that the leaching efficiency of lithium increases firstly and then decreases with the increase of H2SO4 concentration, roasting temperature and roasting time, more than 91 % lithium are leached under optimized experimental conditions: 50 wt% H2SO4, roasting temperature of 650°C, and roasting time of 2 h. Meanwhile, the selectivity of lithium is up to 88 % which means overwhelming majority lithium have been separated completely. In addition, the factors which cause sintering phenomenon such as high temperature and longtime roasting will change the microstructure of roasted product extremely, which is inconducive to the subsequent water leaching. The investigations of the phase conversion mechanism by thermodynamic simulation and instrumental characterization indicated that the leaching behaviors of valuable metals might be influenced by the different reaction pathways of the compounds while changing roasting conditions. The recovery process can be divided into two stages: the partial structure of LiNixCoyMnzO2 was destroyed to form metal sulfates at ambient temperature and the solid phase reaction between unreacted LiNixCoyMnzO2 which contained Li+ in the unstable layered structure with transition metal sulfates at high temperature. The fully reduced low-valence states are mainly (NiO)0.75(MnO)0.25 and CoO, and basically all Li within the crystalline LiNixCoyMnzO2 was de-intercalated and transformed to soluble Li2SO4. These results further confirm the advantages of H2SO4 roasting on selective Li recovery, and provide a better understanding of conversion mechanism by combining theory and experiments.

Rapid precipitation of vanadium from solution using melamine as precipitant with high efficiency

AbstractBACKGROUNDSelective catalytic reduction (SCR) can effectively remove NOx from flue gas in a coal‐fired power plant. As the core of SCR technology, the catalyst has a limited lifespan. The process of reducing acid leaching and roasting water leaching can efficiently extract vanadium and tungsten from spent SCR catalyst. For vanadium recovery in solution, the traditional vanadium precipitation process by NH4Cl has the disadvantages of consuming high precipitant dosage and producing large amounts of waste water.RESULTSThis study focused on the vanadium precipitation process using a solution containing vanadium obtained from adopting some steps to handle the spent SCR catalyst. First, melamine was screened out as the vanadium precipitant. Then, the vanadium precipitation conditions were optimized as follows: solution pH value of 1.0, n(C3H6N6)/n(V) = 0.8, 110 °C and 30 min. Under the best precipitation conditions, the vanadium precipitation efficiency of melamine can reach 99.32%. Finally, the mechanisms of vanadium precipitation by NH4Cl and melamine were discussed, respectively.CONCLUSIONThe results showed that the vanadium precipitation process by NH4Cl followed the chemical reactions between NH4+ and vanadium ions, while the vanadium precipitation process by melamine followed the redox and complexation reactions which essentially belonged to the types of chemical adsorption. The vanadium precipitation product was roasted at 550 °C for 2 h to obtain the vanadium product. The main component of the vanadium product was V2O5, with a purity of 93.27 wt%. The total recovery efficiency of vanadium from spent SCR catalyst was 78.24%. © 2024 Society of Chemical Industry (SCI).

Recovery of Li from waste Li-containing Al electrolytes with high F and Na contents by using a CaO roasting–water leaching process

With the increasing demand for Li, the recovery of Li from solid waste, such as Li-containing Al electrolytes, is receiving growing attention. However, Li-containing Al electrolytes often contain large amounts of F, leading to environmental pollution. Herein, a new method for preparing water-soluble Li salt from waste Li-containing Al electrolytes with high F and Na contents is proposed based on CaO roasting and water leaching. The effects of different roasting and leaching conditions on the Li leaching efficiency and reaction pathway were systematically investigated. Under the optimum processing conditions, the Li leaching efficiency reached 98%, while those of Na and F were 98.41% and 0.24%, respectively. Phase evolution analysis showed that the addition of CaO promoted the conversion of LiF and Na2LiAlF6 to Li2O, whereas F entered the slag phase as CaF2, which could be reused as a raw material for steel refinement. Overall, this study proposes an efficient and environmentally friendly method for the treatment and resource utilization of waste Al electrolytes with high F and Na contents.

Comprehensive recovery of valuable metals from spent LiCoO2 cathode material by a method of NH4HSO4 roasting- (NH₄)₂S leaching

With the retirement of rechargeable lithium-ion batteries (LIBs) entering peak period, the recycling of LiCoO2 has become a hot topic. The spent LiCoO2 batteries are rich in carcinogenic metal Co and high-value metal Li, so they are needed to be recovered in time before causing serious environmental pollution and resource waste. Thus, a clean and efficient process is proposed for the comprehensive recovery of Li, Co, and Mn from the spent LiCoO2 cathode material. It is an integrated pyro-hydrometallurgical process, mainly consisting of sulfation roasting with NH4HSO4 and selective sulfurizing leaching with (NH4)2S solution. Using low melting point ammonium salt as roasting auxiliary, 99.7 %, 99.1 %, and 99.8 % of Li, Co, and Mn in the spent LiCoO2 powder can be converted into sulfate when roasting under 400 ℃ for 75 min. Then 99.5 % of Li in the sulfation roasting product can be selectively leached with an aqueous (NH4)2S solution, which is also a precipitator for Co, Mn, and impurity Al. In addition, 96.4 % of Co and 96.0 % of Mn in the sulfurized precipitate can be separated with impurity elements Al and C via oxidation roasting–water leaching process. Compared to traditional recycling, this method can achieve a higher Li leaching rate at a low roasting temperature, and the recovery of Co and Mn can be realized without adding any acid/base reagents or impurity cations. More importantly, the circulation of NH3/NH4+ and S element in the recycling system has been taken into account, making it an environmental-friendly and economical process. In this study, the sulfation reduction roasting and selective sulfurizing leaching mechanisms were both revealed in details, and the behavior of the valuable elements was also elucidated.

Stepwise extraction of lithium and potassium and recovery of fluoride from aluminum electrolyte wastes through calcification roasting-two-stage leaching.

Aluminum electrolyte is a necessity for aluminum reduction cells; however, its stock is rising every year due to several factors, resulting in the accumulation of solid waste. Currently, it has become a favorable material for the resources of lithium, potassium, and fluoride. In this study, the calcification roasting-two-stage leaching process was introduced to extract lithium and potassium separately from aluminum electrolyte wastes, and the fluoride in the form of CaF2 was recycled. The separation behaviors of lithium and potassium under different conditions were investigated systematically. XRD and SEM-EDS were used to elucidate the phase evolution of the whole process. During calcification roasting-water leaching, the extraction efficiency of potassium was 98.7% under the most suitable roasting parameters, at which the lithium extraction efficiency was 6.6%. The mechanism analysis indicates that CaO combines with fluoride to form CaF2, while Li-containing and K-containing fluorides were transformed into water-insoluble LiAlO2 phase and water-soluble KAlO2 phase, respectively, thereby achieving the separation of two elements by water leaching. In the second acid-leaching stage, the extraction efficiency of lithium was 98.8% from water-leached residue under the most suitable leaching conditions, and CaF2 was obtained with a purity of 98.1%. The present process can provide an environmentally friendly and promising method to recycle aluminum electrolyte wastes and achieve resource utilization.

Recovery of rare earth elements from sedimentary rare earth ore via sulfuric acid roasting and water leaching

Rare earth elements were extracted using a sulfuric acid roasting-water leaching process. The effect of acid roasting on a new type of low-grade sedimentary rare earth ore found in Guizhou Province, China was analyzed using X-ray diffraction and scanning electron microscopy. A systematic study was conducted on process parameters such as amount of acid, roasting temperature, roasting time, water leaching temperature, and leaching time. The results reveal that the total recovery of rare earth elements reaches 81.37%, which is 3.1 times higher than that achieved through direct acid leaching, under the optimal conditions. In addition, the leaching rate of heavy rare earth elements reaches 72.53%. Rare earth elements and some other valuable metals are transformed into soluble sulfate through the local decomposition of clay minerals under the action of the sulfuric acid attack. The dissolution rates of aluminum, iron, and titanium ions are 34.94%, 17.05%, and 62.77%, respectively. The precipitation rate of Ti reaches 99%, and the loss of rare earth ions in the solution is less than 1%. Meanwhile, the results of a leaching kinetics analysis indicate that the leaching process of rare ions is controlled by diffusion. Precious metal ions such as iron and aluminum in the leaching solution can reduce the adsorption of rare earth ions by kaolinite. This study efficiently recovered rare earth ions under conditions of low calcination temperature and direct water leaching, resulting in reduced energy consumption of the extraction process and acidity of the leaching solution. These findings provide a solid foundation for the further separation and extraction of rare earth ions.

Thermal decomposition behavior of Mountain Pass rare earth concentrate in air/CO2 atmosphere

The Mountain Pass mine is recognized as one of the world’s primary sources of rare earth minerals. These rare earth minerals mainly consist of bastnaesite and a small amount of monazite phosphate, which cannot be decomposed and recovered through conventional oxidative roasting and hydrochloric acid leaching process. An efficient, clean, and economical process called the "combined method" was proposed for the utilization of the Mountain Pass mine to extract rare earths from Mountain Pass rare earth concentrate (MPREC). The main steps of this process include weak oxidation atmosphere roasting, step leaching of hydrochloric acid, solid-liquid separation, the monazite slag with sulfuric acid roasting water leaching, etc. In this paper, the roasting process of MPREC under a weak oxidation atmosphere was investigated. The study examines the thermal decomposition kinetics, phase transition process, and leaching behavior of MPREC in air/CO2 atmosphere. Results show that, the activation energy (Ea) for MPREC thermal decomposition in air and CO2 atmosphere are 146 and 320 kJ/mol, respectively. At temperature above 500 °C in air or above 700 °C in CO2 atmosphere, REOF are generated from bastnaesite through an in-situ reaction with CaO, which is decomposed from CaCO3, to form CaF2 and rare earth oxide (REO). Thus, F is regulated into solid phase. In an oxidizing atmosphere, the thermal decomposition of bastnaesite is accompanied by the rapid oxidation of Ce(Ⅲ). In contrast, the oxidation of Ce(Ⅲ) in a CO2 atmosphere is significantly inhibited. At 700 °C, the oxidation rate of Ce in air is 74.09%, while in a CO2 atmosphere, it is only 33.83%. The hydrochloric acid leaching experiment shows that, the leaching rate of rare earth after roasting at 600 °C under an air atmosphere reaches to 82.9%, and it reaches 87% after roasting at 800 °C under a CO2 atmosphere. The reduction of Ce oxidation in a weak oxidizing atmosphere significantly improves the leaching rate of Ce.

Preferentially selective recovery of lithium from spent LiCoO2 by sulfation roasting of MnSO4

The selective recovery of valuable metals from spent lithium-ion batteries (LIBs) has engrossed significant scientific attention due to prompting environmental potential and addressing resource scarcity. Hereby, we propose an innovative and economically more environment friendly manganese sulfate roasting water leaching process to selectively recuperate lithium preferentially from spent LiCoO2 batteries. Notably, it is found that manganese sulfate roasting agents can be recirculated thoroughly without SOx emission during the sulfation roasting process. To better understand the conversion mechanism of LiCoO2, the effects of numerous parameters including, roasting time, roasting temperature, and a molar ratio of MnSO4/LiCoO2 during the recovering process were investigated. An exceptional recovery rate of 98.62% with a remarkable lithium selectivity of 99.35% by manganese sulfate leaching was achieved. This work proposes an environmentally friendly and promising strategy to extract lithium selectively from spent lithium cathodes without introducing SOx emission.

Directly prepared Al-doped FeMoO4 materials from spent hydrogenation catalysts via a low-temperature green process

Environmental-friendly recovery of critical metals from solid wastes directly for new energy materials received lots of concerns. In this paper, the FeMoO4(FMO) materials for new energy storage materials are directly prepared from spent hydrogenation catalysts by “low temperature sodium roasting-water leaching” process. The leaching rates after roasting of main metal element Mo, Al and Ni from wastes are: 91.8 %, 1.7 % and 0.1 %, respectively, then the leachate was directly generated Al-doped FMO materials (Mo/Al = 50:1). As for the application of lithium ion battery anode, the specific capacity is 363.5 mAh/g after 200 cycles, 55 % improvement compared with bare FMO. The tunable phase transformation by roasting temperature regulation is the core to achieve the selective water leaching of Mo, and the intrinsic mechanism have been investigated by calculation and experiments. Finally, the economic evaluation of the whole process shows that for every 1 kg of spent catalyst processed and made into FMO, a profit of $20.8 can be obtained, indicating promising economic benefits. This process provides an environmental technology route for the development of new energy industries under carbon neutrality goals.

An innovative method for the efficient and selective extraction of Co, Ni, Cu, and Mn from oceanic cobalt-rich crusts by ammonium sulfate roasting: Behavior, roasting kinetics and mechanism

The exploitation of deep-sea mineral resources is an unavoidable choice for sustainable human development, given the diminishing or even depletion of land-based mineral resources and the increasing demand for critical metals in global industrialization. This work presents a new method for the selective extraction of Co, Ni, Cu, and Mn from oceanic cobalt-rich crusts through the ammonium sulfate roasting-water leaching process. The metal oxides in oceanic cobalt-rich crusts were transformed to sulfates or ammonium sulfate salts by sulfation roasting and then by selective decomposition roasting to transform Fe2(SO4)3 to Fe2O3, while the other metals remained in sulfate forms. Under the sulfation roasting temperature of 450 ℃ for 120 min and the mass ratio of (NH4)2SO4 to cobalt-rich crusts of 1.2:1, the decomposition roasting temperature of 650 ℃ for 90 min and the air flow rate of 2 L/min, the leaching efficiencies of Mn, Co, Ni, and Cu reached 98.3 %, 93.7 %, 88.7 %, and 66.3 %, while the Fe leaching efficiency was only 0.9 %. The analysis of the mechanism shows that the Mn in oceanic cobalt-rich crusts first formed (NH4)2Mn2(SO4)3 during roasting and then decomposed to MnSO4; Co, Ni, and Cu first formed (NH4)2Co2(SO4)3, (NH4)2Ni2(SO4)3, (NH4)2Cu3(SO4)4 and then decomposed to their respective sulfates; Fe first formed (NH4)3Fe(SO4)3 and finally decomposed to Fe2O3. Ammonia released during the roasting process can be absorbed with dilute sulfuric acid and then regenerated (NH4)2SO4 with an ammonia recovery of 87.3 %. The process proposed in this study is not only efficient, clean, and selective, but also achieves full component recovery from oceanic cobalt-rich crusts, with obvious environmental and economic advantages over the traditional reduction roasting-acid/ammonia leaching.

Efficient separation and extraction of lithium from low-grade claystone by chloride salt-enhanced roasting process

The exploitation and extraction of lithium resources are mainly hard rock ore and salt lake brine. However, with the increasing demand of lithium resource market, the developed resources cannot meet the demand gap of lithium resource. In China, as a natural lithium source with considerable reserves, the exploitation and extraction of clay lithium resources can effectively relieve the pressure of lithium resource import. In this study, a green extraction process of low-grade lithium claystone treated by chlorination roasting-water leaching was proposed. The mechanism of lithium conversion under chlorination conditions was also investigated to reveal the fugitive form of lithium in claystone and its migration behavior during the extraction process. The experimental results showed that the optimal reaction conditions were 800 °C, the mixing ratio of the claystone and CaCl2 was 1:0.5, the roasting time was 2 h, the leaching temperature was 90 °C, and the leaching time was 3 h. Under such treatment conditions, the recovery efficiency of lithium could reach 92.35 %. Compared with the traditional roasting-acid leaching method of lithium claystone, it is more environmentally friendly and economical, and has good application prospects.

Selective recovery of vanadium from high-chromium vanadium slag by a mechanically activated low-sodium salt roasting-water leaching process

Stabilizing the toxic element chromium and improving vanadium recovery are crucial for the economic and efficient treatment of high chromium-bearing vanadium slag. This study employed mechanical activation to enhance vanadium extraction from high chromium-bearing vanadium slag through a low-sodium salt roasting-water leaching process. The effects of Na2CO3 addition, roasting temperature, and roasting time were investigated to determine the optimal roasting parameters. Subsequently, the effects of mechanical activation on the morphological and microstructural changes were characterized using a laser diffraction particle size analyzer, a micropore physisorption analyzer, scanning electron microscopy, energy dispersive spectrometry, and X-ray diffraction. The effects of mechanical activation on the extraction of vanadium and chromium were determined by leaching the roasted samples after varying activation times. The results indicated that mechanical activation decreased particle size, increased surface area, and enhanced the activity of the samples. The leaching results demonstrated that mechanical activation enhanced the leaching efficiency of vanadium by 5 %, reduced the optimum roasting temperature by 50 ℃, and slightly increased the leaching efficiency of chromium by 0.5–1 %.

One-step extraction of zinc and separation of iron from hazardous electric arc furnace dust via sulphating roasting-water leaching

The electric arc furnace dust (EAFD) is rich in zinc and iron, which is classified as a hazardous solid waste and generated during steelmaking in electric arc furnace. As an important secondary resource for zinc, the zinc extraction from EAFD has become crucial resource utilization direction. In this work, a novel process is proposed to extract zinc and separate iron for one-step through co-roasting EAFD with NH4HSO4, followed by water leaching. Through the co-roasting of EAFD and NH4HSO4, the zinc-containing phases in EAFD were transformed into ZnSO4, and the iron-containing phases were transformed into Fe2O3. The separation of Zn and Fe was then performed by water leaching. The determined optimal experimental conditions were the mass ratio of NH4HSO4/EAFD of 0.75, the roasting temperature of 600 ℃, the holding time of 3 h, and the leaching time of 1 h, in which 91.14% of Zn and 3.25% of Fe were dissolved in the leaching solution. Afterwards, the reaction mechanisms of the co-roasting process of EAFD and NH4HSO4 were revealed. When the temperature increased, the zinc-containing phase (NH4)2Zn(SO4)2 and iron-containing phase NH4Fe(SO4)2 were first generated. (NH4)2Zn(SO4)2 was then successively converted to ZnSO4, and NH4Fe(SO4)2 was decomposed into Fe2O3 and Fe2(SO4)3. SO3 produced by NH4HSO4 decomposition in the roasting process can continue to participate in the sulfation process. Subsequently, zinc was enriched in the leaching solution, which can be considered as the raw material for Zn electrodeposition. Iron entered the leaching residue as Fe2O3, which can be used as the raw material for ironmaking. The result of leaching toxicity test showed that most of the heavy metals (i.e., Pb and Cr) were fixed in the leaching residue, which met the national standards for leaching toxicity. This study provides a theoretical reference and technical support for zinc extraction from zinc-containing solid wastes.

An efficient and clean method for the selective extraction and recovery of manganese from pyrolusite using ammonium sulfate roasting-water leaching and carbonate precipitation

Manganese is a fundamental metal raw material that plays a crucial role in sustainable economic and social development. With the battery industry undergoing rapid advancements in the recent years, there has been a consistent rise in the demand for manganese. This study proposes a new scheme for selectively extracting manganese from pyrolusite by the sulfation roasting-water leaching process, while the iron is enriched in the leaching residues in the form of Fe2O3. With a 1.5:1 mass ratio of ammonium sulfate to pyrolusite, a roasting temperature of 650 °C, and a duration of 120 min, the extraction efficiency of manganese was observed to be 96.15%, while iron extraction was only 1.45%. And the leaching residues containing mainly 54.62% SiO2 and 36.10% Fe2O3 were obtained, which could be enriched with iron by flotation. Subsequently, the purified leaching solution was subjected to manganese precipitation using ammonium bicarbonate. Under optimal conditions, the precipitation efficiency of manganese was 99.63%, and the manganese carbonate products were obtained with a particle size of approximately 1–1.5 μm spherical structure, which met the requirements of the industrial manganese carbonate quality standard (HG/T4203-2011). The mechanism studies have proven that MnO2 in pyrolusite was first converted to (NH4)2Mn2(SO4)3 and then to MnSO4, while Fe2O3 was converted to (NH4)3Fe(SO4)3, NH4Fe(SO4)2, Fe2(SO4)3 in turn and finally decomposed to Fe2O3. The ammonia produced during the roasting process can be absorbed to regenerate ammonium sulfate products, and the filtrate from the manganese precipitation can also be crystallized by evaporation to obtain ammonium sulfate products, both of which can be used in the roasting process to reduce production costs. The process of this study proposed is simple and does not require any strong acid or alkali reagents. It offers high efficiency, is environmentally friendly, and is easy to operate, thus providing a new idea and scheme for the multi-component resource utilization of manganese oxide ore.

Thermal treatment of K-feldspar for potassium recovery using NaCl–CaCl2 mixture and its mechanism study

The increasing global demand for potassic fertilizer, due to the population growth, indirectly forces many countries to search for an alternative source of potassium. The present study reports the recovery of potassium from K-feldspar (K2O 12.5 wt. %) through the chlorination roasting—water leaching process applying a mixture of CaCl2 and NaCl. Individual use of NaCl and CaCl2 results in the extraction of ∼50% and 92% of potassium, respectively. Though the K-extraction is very high via CaCl2 roasting process, it adds an extra cost to the process as compared with NaCl. However, the use of salt mixture CaCl2: NaCl (40:60 wt. %) recovers potassium to the tune of ∼98% at optimum conditions (850 °C, 60 min roasting period, salt mixture: feldspar 1:1). Various experimental parameters such as temperature, time, and salt to feldspar ratio were optimized to recover maximum potassium from the complex structure of K-feldspar. The experimental results are explained using a theoretical extraction model, further supported by characterizations of the samples using XRD, SEM-EDX and FTIR studies.

High-Selective Lithium Extraction from Spent LiFePO4 by Battery Roasting-Water Leaching Method

High-Selective Lithium Extraction from Spent LiFePO4 by Battery Roasting-Water Leaching Method

Recovery of rare earths and niobium from a potash feldspar ore beneficiation tailings

In this paper, a hydrometallurgy process was investigated for recovering rare earths (REs) and niobium (Nb) from a potash feldspar ore beneficiation tailings. Firstly, the REs and Nb could be leached from potash feldspar ore beneficiation tailings by sulfuric acid roasting-water leaching process, and the effects of the roasting temperature, roasting time, sulfuric acid addition, leaching temperature, ratio of solid to liquid and leaching time for REs and Nb leaching rate were investigated, and the REs and Nb leaching rate could reach to 89.9% and 80.6% respectively at the optimum test conditions. Then, the REs was removed from the leaching liquor by adding HF, the precipitation rate could reach to 91.53%. The rare earths fluoride (REF3) with the purity 60.51% was obtained, and the REs recovery rate reached to 70.6%. Finally, the Nb was recovered by solvent extraction with methyl isobutyl ketone (MIBK), Nb extraction rate could be reached 99.59%. The loaded organic phase could be stripped easily, nearly all niobium was stripped using 0.5 M sulfuric acid as strip liquor. The Nb2O5 products with the Nb2O5 content 87.61% was obtained by ammonia precipitation-roasting process, and the recovery Nb was 76.31% in the whole process.

Extraction efficiency of metals from low-nickel matte via NH4Cl roasting-water leaching process and synthesis of (Ni,Cu,Co)Fe2O4 photocatalyst

Extraction efficiency of metals from low-nickel matte via NH4Cl roasting-water leaching process and synthesis of (Ni,Cu,Co)Fe2O4 photocatalyst

Extraction of Metals from Nickel Concentrate by Low-Temperature NH4HSO4 Roasting-Water Leaching and Efficient Separation of Iron and Copper

Extraction of Metals from Nickel Concentrate by Low-Temperature NH4HSO4 Roasting-Water Leaching and Efficient Separation of Iron and Copper