Sulfation Roasting 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.

Synchronous conversion and targeted separation of valuable metals from spent hydroprocessing catalysts

Sustainable recovery of spent hydroprocessing catalysts (sHPC) is crucial to address environmental and resources issues. However, the existing recycling methods recover limited metal components due to uncontrollable phase transition and complicated separation procedures, leading to poor utilization efficiency of the sHPC. Herein, a synchronous conversion and targeted separation strategy was designed to achieve full-component recovery of the sHPC. The insoluble phases of Al/Ni/Mo/V could be completely transformed into soluble sulfate via sulfation roasting at 400 °C. The roasted sulfate products can be directly leached by water with efficiencies close to 100%. Targeted separation of the multi-ions could be achieved based on their species difference in the leachates. Anionic HV10O285− and HMoO4− species were co-extracted by basic trioctylamine extractant, while cationic Ni2+ and Al3+ were extracted by acidic HBL110 and P204, respectively. Benefiting from the sulfation roasting and selective extraction procedures, the recovery efficiencies of Al, Ni, Mo, and V elements achieved 97.2%, 95.8%, 94.1%, and 94.3%, respectively. The explored synchronous conversion and targeted separation strategy is an industrial-feasible candidate for sustainable utilization of the retired catalysts.

Extraction of Rare Earth Metals from NiMH Battery Scrap via Selective Sulfation Roasting

Extraction of Rare Earth Metals from NiMH Battery Scrap via Selective Sulfation Roasting

Influence of SiO2 and MgO on Extraction of Nickel and Cobalt from Laterite Ores by Sulfation Roasting Leaching Process

Influence of SiO2 and MgO on Extraction of Nickel and Cobalt from Laterite Ores by Sulfation Roasting Leaching Process

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.

Simultaneous extraction of metals from nickel concentrate via NH4HSO4 roasting−water leaching process and transformation of mineral phase

Low-temperature NH4HSO4 roasting−water leaching process was proposed to extract Ni, Cu and Co from Jinchuan nickel concentrate simultaneously. The effects of some factors on the metal extraction in the process of sulfation roasting and water leaching were systematically investigated. The results showed that 95.7% Ni, 98.9% Cu and 96.8% Co were extracted under optimized roasting and water leaching conditions. The transformation evolution of mineral phases was determined through the means of DTA−TG, XRD, SEM and EDS. The valuable metals were effectively extracted by being transformed to water-soluble metal−ammonium complexes under the sulfation reaction of NH4HSO4 and decomposed SO3 and NH3. Iron impurities in the leach solution can be removed by elevating leaching temperature to combine iron ions with ammonium ions to form NH4[Fe3(SO4)2(OH)6] precipitate. The volatiles collected in the roasting process were determined to be (NH4)2SO4 and (NH4)2SO3, which can be used as reagents for sulfation roasting.

Optimization of sulfation roasting and water leaching for zinc recovery from zinc waste

Çinko atıkları çinko üretimi için ekonomik değere sahip bir kaynaktır. Ancak, aynı zamanda içerdiği ağır metaller nedeniyle tehlikeli atık olarak da sınıflandırılırlar. Bu çalışmada sülfatlayıcı kavurma ve su liçi koşullarının metal çözünürlüğüne (Zn, Pb, Fe ve Ca) etkisi araştırılmıştır. Çeşitli sıcaklarda ve sürelerde amonyum sülfatla yapılan kavurma testleri sonrasında, kavrulmuş her bir numunenin mineralojik yapısı incelenmiştir. Kimyasal analiz sonuçlarına göre atığın %16,6 Pb, %12,5 Zn ve %7,5 Fe içerdiği tespit edilmiştir. Amonyum sülfatla ile kavurma işlemi sonunda, atığın sahip olduğu ZnFe2O4 ve Zn2SiO4 yapılarının çinko sülfata (ZnSO4) dönüştüğü saptanmıştır. Proses ekonomisi ve enerji tasarrufu göz önüne alınarak, yaklaşık %91 oranında çinko geri kazanımı için optimum çalışma şartı, 1:2 (g:g) oranında atık/amonyum sülfat oranı, 625 ºC kavurma sıcaklığı, 4 saat kavurma süresi, 25 ºC liç sıcaklığı ve 2 saat liç süresi olarak belirlenmiştir. Bu şartlar altında kurşun çözünmediği, kalsiyum ve demirin çözünürlüklerinin sırasıyla %41,2 ve %16,8 olarak tespit edilmiştir.

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.

Influence of Alkali Salts on Extraction of Nickel From Serpentine-Rich Ore Through Sulfation Roasting Leaching Process

Influence of Alkali Salts on Extraction of Nickel From Serpentine-Rich Ore Through Sulfation Roasting Leaching Process

The recovery of valuable components from secondary lead dust by a multi-step hydrometallurgical process

Multicomponent secondary lead dust (MSLD) is a significant resource that contains valuable metals such as lead, zinc, and cadmium, along with potentially harmful components such as salts (K, Cl) and As. In this study, an economic and efficient method was proposed for the recovery of MSLD, including desalination, sulfation roasting, water leaching, and zinc plate cementation. Firstly, desalination was conducted to remove K and Cl, and 90.4% of K and 76.5% of Cl were selectively removed with a low NaOH consumption of 1.00 g/t dust at 40 °C and pH 7.5. Subsequently, the desalted residue, enriched with Cd, Zn, As, and Pb, was treated by H2SO4 roasting, which facilitated the transformation of arsenites and sulfides into As2O3 and sulphates. The recovery of As was achieved with a volatilisation efficiency of 93.2% as As2O3 using 65% H2SO4 at 350 °C. Finally, the recovery of Cd was accomplished through zinc plate cementation, resulting in the cementation of 99.4% Cd as Cd sponge with a purity of 96.3%. The residue from water leaching and cemented solution gathered 98.2% Pb and 94.6% Zn. The multi-step process realised the selective removal of K and Cl salts as well as the comprehensive recovery of valuable elements.

The Recovery of TiO2 from Ilmenite Ore by Ammonium Sulfate Roasting–Leaching Process

TiO2 production is a key part of Ti metallurgy and Ti recycling, and the process itself has turned out to be energy-consuming and material-consuming. New technologies are needed to utilize complex Ti ores, such as ilmenite, and reduce the carbon footprint of TiO2 extraction. Ammonium sulfate roasting has been revealed as an efficient way to carry out phase transformations of complex minerals. A low-temperature sulfation roasting approach was studied to chemically breaking down the crystal structure of ilmenite and generate metal soluble sulfates simultaneously. These roasted products were introduced to water leaching, then the residue of the water leaching was leached by diluted HCl acid, and the TiO2 product was enriched in the leaching residue. The effects of roasting temperature, roasting time, ilmenite-to-ammonium sulfate mass ratio, ilmenite particle size, and second-stage roasting on iron removal and titanium loss leaching efficiency were systematically studied. The results show that the optimum roasting conditions were a roasting temperature of 500 °C, a roasting time of 210 min, an ilmenite-to-(NH4)2SO4 mass ratio of 1:7, and an ilmenite particle size of below 43 µm. Under optimized conditions, the TiO2 grade in the obtained synthetic rutile reached 75.83 wt.%. Furthermore, the phase transformation and reaction mechanism during roasting are discussed and interpreted.

Study on the synergistic mechanism of sulfuric acid and ammonium sulfate on the extraction of metals from ferronickel slag by roasting

Due to the complex occurrence state of valuable metals in ferronickel slag, the metal extraction rate of the ammonium sulfate roasting process is insufficient. This study proposes a more cost-effective strategy to extract ferronickel slag valuable metals by mixed roasting with sulfuric acid and ammonium sulfate. The effects of the reagent dosage, roasting temperature and roasting time on the metal extraction rate of (NH4)2SO4 roasting, H2SO4 roasting and (NH4)2SO4 and H2SO4 mixed roasting were compared and analyzed. XRD, SEM and TG-DSC were used to reveal the mechanism of mixed roasting to optimize the metal extraction effect from ferronickel slag. The results showed that mixed roasting with a mass ratio of ammonium sulfate to sulfuric acid of 1:1 could reduce the dosage of ammonium sulfate roasting by 2.5 times, and the leaching rates of valuable metals Mg, Fe, Al and Cr increased by 17.46%, 42.8%, 26.06% and 24.89%, respectively. Likewise, comparing to sulfuric acid roasting, the acid consumption of mixed roasting was reduced by two-thirds, and the metal leaching rate was increased by 10.08%, 12.98%, 18.76% and 45.18%, respectively. The “synergistic” mechanism of sulfuric acid and ammonium sulfate on metal extraction by roasting is as follows: ① Before roasting, sulfuric acid partially destroys the surface structure of ferronickel slag particles, forming more pores and channels, creating conditions for the rapid diffusion of ammonium sulfate and sulfuric acid into the slag particles. ② Sulfuric acid reacts with ammonium sulfate to form NH4HSO4, which is more acidic and can accelerate conversion to soluble metal complexes. ③ The addition of sulfuric acid greatly advances the initial temperature of the reaction between ammonium sulfate and metals, widening the temperature range of sulfation roasting and reducing the difficulty of the reaction between ammonium sulfate and metals in ferronickel slag. ④ The liquid–solid reaction increases the reaction interface, improves the mass and heat transfer efficiency and enhances the degree of metal acidification.

Selective extraction of rare earth elements from florencite-rich sedimentary rare earth ore by sulfation

Simultaneous recovery of lanthanum (La), cerium (Ce), and niobium (Nd) from sedimentary rare earth ores (SREO) has been carried out by a hydrometallurgical process using a sulfation roasting - leaching process (SRLP). Response surface methodology (RSM) was used to optimize the process parameters. The optimum condition for rare earth elements (REE) recovery was determined to be roasting temperature 700 °C, roasting liquid/solid ratio (L/S) 1.35 mL/g, roasting time 15 min, leaching temperature 90 °C, leaching L/S 20 mL/g, and leaching time 4.24 h. The maximum predicted recoveries for La, Ce, and Nd were 80.93 %, 74.48 %, and 97.49 %, respectively. The leaching verification experiments showed that the % leaching under the optimal conditions were La 88.02 %, Ce 78.87 %, and Nd 89.75 %, respectively. The % leaching relative deviation between the experimental and predicted values was less than 10 %, confirming the reliability of the experiments. Mechanism analysis shows that after SRLP, the florencite and kaolinite in SREO are more likely to be selectively converted to aqueous solution by H2SO4, thus achieving the separation of La, Ce, and Nd from SREO. This provides a valuable reference for the recovery of REE from florencite-base rare earth ore.

Selective recovery of lithium from spent lithium-ion battery by an emission-free sulfation roasting strategy

As the economy recovered from the COVID-19 epidemic, the price of Li2CO3 skyrocketed to the highest. Recovery of lithium from spent lithium-ion batteries (LIBs) is significant for addressing lithium shortage and environmental issues. Sulfation roasting is often accused of being unsustainable and not environmentally friendly due to the consumption of expensive sulfation reagents and emission of SO2. Herein, a novel and green recycling process for selective separation of lithium from spent LiMn2O4 (LMO) batteries was proposed based on a SO2 emission free sulfation roasting with waste copperas. Lithium in the cathode power were selectively sulfated into soluble Li2SO4 with a conversion of 99.50%, while manganese was riched as insoluble oxides. At low temperature (e.g. 400 °C), LMO were difficult to be sulfated due to the limited sulfation ability of SO42- and mass transfer. When the roasting temperature reached 700 °C, SO2 was produced and the generated SO2 reacted in situ with the LMO to form Li2SO4 and MnSO4, preventing SO2 emissions into the atmosphere. As temperature continued to increase, the MnSO4 acted as a sulfation reagent and sulfated the unreacted LMO. All the sulfur was immigrated in Li2SO4, rather than being emitted as SO2.

High-efficiency and environment-friendly separation and recovery of manganese from braunite via the ammonium sulfate roasting-water leaching process: Behavior and mechanism

Manganese ore is a key raw resource for the development of new energy, new materials, and other emerging industries, as well as a basic raw material for the steel industry. As resources of the main high-grade pyrolusite are depleting, it is of strategic importance to develop and utilize other manganese oxide ores. This paper proposed the two-stage ammonium sulfate roasting-water leaching process for high-efficiency and environment-friendly separation and recovery of manganese from braunite ((Mn2O3)3MnSiO3). The first-stage was sulfation roasting, which could convert Mn and Fe in braunite into (NH4)2Mn2(SO4)3, (NH4)3Fe(SO4)3, NH4Fe(SO4)2 and Fe2(SO4)3; the second-stage was selective roasting, that could decompose (NH4)2Mn2(SO4)3 into MnSO4, and decompose (NH4)3Fe(SO4)3, NH4Fe(SO4)2 and Fe2(SO4)3 into Fe2O3. Under the optimum leaching conditions, the leaching efficiency of Mn reached 97.79%, while the leaching efficiency of Fe was only 1.52%, and a solution with a concentration of 93.36 g/L for Mn was obtained. The NH3 volatilized during roasting was absorbed by the water solution to get an ammonia solution which was used to precipitate Mn from the leaching solution. The precipitation efficiency of Mn was 99.48% under optimal conditions. Manganese oxide products such as Mn3O4 or Mn2O3 could be obtained by drying and calcining Mn-rich precipitates. The final filtrate was neutralized with H2SO4 and then evaporated and crystallized to give an (NH4)2SO4 product, enabling the recycling of ammonia. The process proposed in this paper is not limited to the treatment of braunite, but can also be applied to the processing of other manganese oxide ores and has a wide application potential. Compared to the traditional reduction roasting-acid leaching process, this process is more efficient and environmentally friendly.

Influence of calcination temperatures on lithium deportment by screening hard rock lithium

Calcination of spodumene is a pre-treatment stage in preparation for sulfation roasting and leaching in lithium recovery. During calcination, α-spodumene (less reactive, monoclinic crystal structure) is converted to β-spodumene (more reactive, tetragonal crystal structure). A third, metastable γ-phase has been identified at lower temperatures than full conversion to the β-phase. It has been previously observed that calcination greatly alters the physical properties of the various minerals in pegmatite ores, impacting comminution energy and liberation. Thus, this work investigates the relationships between calcination temperatures and the physical behaviour of hard rock lithium ores. The results showed that the increase in calcination temperature resulted in a higher lithium deportment in the finest size fraction (−0.6 mm) and thus a higher lithium grade and recovery. The samples calcined at 813.15 K and 1223.15 K did not show a significant increase in lithium grades in the finest size fraction. This work shows the incremental change in the physical properties of various minerals in the ore with increasing calcination temperature.

Selective Sulfation Roasting for Cobalt and Lithium Extraction from Industrial LCO-Rich Spent Black Mass

The extraction of cobalt from secondary resources has become crucial, as cobalt has been identified as a strategically important and critical raw material due to the high risks of supply chain disruptions. In this work, selective sulfation roasting was investigated as a potential recycling strategy to extract cobalt and lithium from an industrial lithium cobalt oxide (LCO)-rich black mass. Additionally, the effect of graphite on metal extraction was studied. In the first set of experiments, the sieved black mass fraction containing both cathode and anode materials was directly roasted in a predetermined composition of gas mixtures of SO2, O2, and Ar for 1 h at 850 °C. The gas composition was determined from Kellogg’s diagram to allow for the selective sulfation of Co and Li. In another set of experiments, the carbon present in the black mass was first removed by roasting the material in Ar for 2 h and then in an Ar and O2 gas mixture for five hours at 600 °C. Afterward, selective sulfation roasting was performed in mixtures of SO2, O2, and Ar gas similar to the previous set of experiments. For comparison, similar experiments were performed at 800 °C. The sulfation roasted black mass was leached in water to study the efficiency of Co extraction into the solution. Interestingly, the presence of carbon was found to be beneficial for Co extraction. The extraction efficiency for the first case (with carbon present in the raw material) was observed to be more than three times higher than in the second case (with carbon removed) for sulfation at 850 °C. The extraction efficiency and purity of the extracted Co were found to be better for higher temperature sulfation roasting conditions due to faster reaction kinetics. It was also found that almost all of the Li could be recovered while extracting Co. The maximum efficiency of the extraction was 99.51% Li and 61.21% Co for roasting under a gas flow of 10% SO2-10% O2-Ar at 850 °C for 60 min. These results suggest that Co and Li can be selectively extracted from the black mass by sulfation roasting pre-treatment followed by leaching in water. In holistic processing, the leach residue can then be further subjected to battery metal processing by state-of-the-art methods.

On the Cobalt Content Upgrade in Nickeliferous Laterites Using Iron (III) Sulfate: A Study Based on Thermodynamics Simulations

Nickel (Ni) and cobalt (Co) are relevant technological metals for the future of the lithium-ion battery (LIB) industry. Based on the current and projected demand for these, an increased interest in developing processing routes to exploit lateritic occurrences has been observed, as these are reported as critical raw materials for future mineral–metallurgical industry. However, the content of Ni and Co in such ores is minimal and requires impracticable mineral-processing operations for concentration before metal extraction. It was identified that information regarding the sulfation roasting of this material is scarce on what concerns the iron sulfates interaction as a function of the temperature. Based on that context, the present work has its purposes associated with the proposition of an alternative chemical pretreatment to upgrade the content of metals of technological interest in lateritic ores through a simple roast–leach process. Thus, the chemical interactions between the mineral sample and iron (III) sulfate (Fe2(SO4)3) through thermodynamic simulations and experimental procedures were explored. The latter included specific water leaching practices for the selective concentration of metals. The equilibrium calculations indicate that Fe2(SO4)3 and FeSO4 tend to decompose at lower temperatures, and considering the higher stability of other metal sulfates, it could be an interesting reagent in this type of process. Regarding the experimental results, the characterization of materials indicates a recovery of Co as high as 73.4 wt.% after sulfation roasting at 500 °C followed by water leaching, with the full content of Iron (Fe) being reported in the insoluble phase. Based on these findings, the present development could be an interesting alternative to consider within operations for the chemical upgrade of cobalt in such types of mineralogical occurrences.

An emission-free controlled potassium pyrosulfate roasting-assisted leaching process for selective lithium recycling from spent Li-ion batteries

Recycling critical metals from spent Li-ion batteries (LIBs) is important for the overall sustainability of future batteries. This study reports an improved sulfation roasting technology to efficiently recycle Li and Co from spent LiCoO2 LIBs using potassium pyrosulfate as sulfurizing reagent. By sulfation roasting, LiCoO2 was converted into water-soluble lithium potassium sulfate and water-insoluble cobalt oxide. Under optimal conditions, 98.51% Li was leached in water, with a selectivity of 99.86%. More importantly, sulfur can be recirculated thoroughly, and the sulfur atomic efficiency can be significantly enhanced by controlling the amount of potassium pyrosulfate. Li ions from the water leaching process were recovered by chemical precipitation. Furthermore, application of this technology to other spent LIBs, such as LiMn2O4 and LiNi0.5Co0.2Mn0.3O2, verified its effectiveness for selective recovery Li. These findings can provide some inspiration for high efficiency and environmentally friendly recovery metal from spent LIBs.

Selective separation and recovery of selenium from copper anode slime by compound leaching followed by sulfate roasting

In this research, selenium was selectively separated and recovered from copper anode slime by a hydrometallurgical-pyrometallurgy union process. In the first compound leaching process with H2SO4, HCl and NaCl media, the pre-removal of As, Sb, and Bi was achieved, and the removal efficiencies were 85 %, 83 %, and 91 %, respectively. Selenium was effectively enriched by approximately 200 % after first compound leaching. Sulfate roasting method was used to selectively separate selenium in the second stage. In this study, more than 99.39 % of selenium was separated with concentrated sulfuric acid amount to first leaching residues ratio (H2SO4/ first compound leaching residues ratio) of 0.30 mL/g at 500 °C for 25 min, which enabled selenium to be selectively separated from the other elements. The kinetics of sulfation roasting of Se-enriched residues show that the selective separation of selenium conforms to the Avrami multiphase liquid–solid region reaction, the activation energy of the reaction is 28.118 kJ/mol, and the reaction efficiencies equation is k = 3.607e-3382/T, indicating that it is internal diffusion-controlled.