Separation of halogens and recovery of heavy metals from secondary copper smelting dust

The separation of halogens and recovery of heavy metals from secondary copper smelting(SCS)dust using a sulfating roasting−water leaching process were investigated.The thermodynamic analysis results confirm the feasibility of the phase transformation to metal sulfates and to gaseous HF and HCl.Under the sulfating roasting conditions of the roasting temperature of 250℃ and the sulfuric acid excess coefficient of 1.8,over 74 wt.%of F and 98 wt.%of Cl were volatilized into flue gas.Approximately 98.6 wt.%of Zn and 96.5 wt.%of Cu in the roasting product were dissolved into the leaching solution after the water leaching process,while the leaching efficiencies of Pb and Sn were only 0.12%and 0.22%,respectively.The mechanism studies indicate the pivotal effect of roasting temperature on the sulphation reactions from various metal species to metal sulfates and the salting out reactions from various metal halides to gaseous hydrogen halides.

Recovery of valuable metals from zinc leaching residue by sulfate roasting and water leaching

Zinc leaching residue(ZLR),produced from traditional zinc hydrometallurgy process,is not only a hazardous waste but also a potential valuable solid.The combination of sulfate roasting and water leaching was employed to recover the valuable metals from ZLR.The ZLR was initially roasted with ferric sulfate at640°C for1h with ferric sulfate/zinc ferrite mole ratio of1.2.In this process,the valuable metals were efficiently transformed into water soluble sulfate,while iron remains as ferric oxide.Thereafter,water leaching was conducted to extract the valuable metals sulfate for recovery.The recovery rates of zinc,manganese,copper,cadmium and iron were92.4%,93.3%,99.3%,91.4%and1.1%,respectively.A leaching toxicity test for ZLR was performed after water leaching.The results indicated that the final residue was effectively detoxified and all of the heavy metal leaching concentrations were under the allowable limit.

Recovery of lead and silver from zinc acid-leaching residue via a sulfation roasting and oxygen-rich chlorination leaching method

A large amount of acid-leaching residue is produced during the conventional Zn hydrometallurgy process,and this residue has a large concentration of a variety of valuable metals.The purpose of this study was to evaluate the ability of a procedure that entails the use of sulfation roasting,water leaching,and chlorination leaching(blowing oxygen technique)to recover Pb and Ag,followed by cooling crystallization and the replacement of Ag with lead sheet,to realize the full recovery of all valuable metals from zinc acid-leaching residue;consequently,good results were achieved.The best results were obtained under the following conditions:a sulfuric acid at 70%of the raw material quality,roasting temperature of 300°C and roasting time of 2 h,followed by the process of leaching the roasted residue for 1 h by applying a water-to-solid ratio of 5꞉1 at room temperature.The recovery rates of Zn and Fe were 98.69%and 92.36%,respectively.The main parameters of the chlorine salt leaching system were as follows:Cl−concentration of 300 g/L,Fe3+concentration of 25 g/L,acid concentration of 2 mol/L,liquid-to-solid ratio of 9 mL:1 g,temperature of 90℃,and leaching time of 0.5 h;this leaching process was followed by filtration separation.These conditions resulted in high extents of leaching for Pb and Ag(i.e.,98.87%and 96.74%,respectively).Finally,the kinetics of the process of Ag leaching using Cl−ions in an oxygen-rich medium was investigated.It was found that the leaching process was controlled by the diffusion of the product layers,and the activation energy was 19.82 kJ/mol.

Zinc extraction from zinc oxidized ore using(NH4)2SO4 roasting−leaching process

An improved method of(NH4)2SO4 roasting followed by water leaching to utilize zinc oxidized ores was studied.The operating parameters were obtained by investigating the effects of the molar ratio of(NH4)2SO4 to zinc,roasting temperature,and holding time on zinc extraction.The roasting process followed the chemical reaction control mechanism with the apparent activation energy value of 41.74 kJ·mol^−1.The transformation of mineral phases in roasting was identified by X-ray diffraction analysis combined with thermogravimetry–differential thermal analysis curves.The water leaching conditions,including the leaching temperature,leaching time,stirring velocity,and liquid-to-solid ratio,were discussed,and the leaching kinetics was studied.The reaction rate was obtained under outer diffusion without product layer control;the values of the apparent activation energy for two stages were 4.12 and 8.19 kJ·mol^−1.The maximum zinc extraction ratio reached 96%while the efficiency of iron extraction was approximately 32%under appropriate conditions.This work offers an effective method for the comprehensive use of zinc oxidized ores.

Lithium and manganese extraction from manganese-rich slag originated from pyrometallurgy of spent lithium-ion battery

Mn and Li were selectively extracted from the manganese-rich slag by sulfation roasting−water leaching.The extraction mechanisms of Mn and Li were investigated by means of XRD,TG−DSC,and SEM−EDS.73.71%Mn and 73.28%Li were leached under optimal experimental conditions:acid concentration of 82 wt.%,acid-to-slag mass ratio of 1.5:1,roasting temperature of 800°C,and roasting time of 2 h.During the roasting process,the manganese-rich slag first reacted with concentrated sulfuric acid,producing MnSO_(4),MnSO_(4)·H_(2)O,Li_(2)Mg(SO_(4))_(2),Al_(2)(SO_(4))_(3),and H_(4)SiO_(4).With the roasting temperature increasing,H_(4)SiO_(4) and Al_(2)(SO_(4))_(3) decomposed successively,resulting in generation of mullite and spinel.The mullite formation aided in decreasing the leaching efficiencies of Al and Si,while increasing the Li leaching efficiency.The formation of spinel,however,decreased the leaching efficiencies of Mn and Li.

An environmentally benign and sustainable process for carbon recovery and efficient defluorination of spent carbon cathode

A systematic and green low-temperature sulfation roasting−water leaching strategy was put forward to achieve a very high fluorine removal rate of 97.82%for spent carbon cathode(SCC),which was believed as a hazardous solid waste.And the carbon could be recycled with a purity of 90.29 wt.%in the flaky microstructure.Thermodynamic analysis and the results of SEM,XRD and EDS indicate that most of the fluoride could convert into water-soluble sulfate at low temperature.And the highest fluorine removal rate could be obtained when<0.15 mm SCC particles were mixed with sulfuric acid at a liquid-to-solid ratio of 1:1,and then roasted at 300℃ for 0.5 h.The sulfate was removed to purify the carbon via water-leaching process.Avrami exponents and corresponding activation energy for the roasting and leaching process demonstrated that both processes are controlled by diffusion.

Decomposition behavior of CaSO_4 during potassium extraction from a potash feldspar-CaSO_4 binary system by calcination

The extraction of potassium from a tablet mixture of K-feldspar ore and CaSO4by roasting was studied with a focus on the effects of the decomposition behavior of CaSO4on the potassium extraction process.The roasted slags were characterized by X-ray diffraction（XRD）,scanning electron microscopy（SEM）,energy-dispersive X-ray spectroscopy,and thermogravimetric（TG）analysis.The XRD analysis revealed that hydrosoluble mischcrystal K2Ca2（SO4）3was obtained by ion exchange of Ca^2＋ in CaSO4 and K^＋ in KAlSi3O8.Meanwhile,the intermediate product,SiO2,separated from KAl Si3O8and reacted with CaSO4to decompose CaSO4.The SEM results showed that some blowholes emerged on the surface of the CaSO4particles when they reacted with SiO2at 1200℃,which indicates that SO2and O2gases were released from CaSO4.The TG curves displayed that pure CaSO4could not be decomposed below 1200℃,while the mixture of K-feldspar ore and CaSO4began to lose weight at 1000℃.The extraction rate of potassium and decomposition rate of CaSO4were 62%and 44%,respectively,at a mass ratio of CaSO4to K-feldspar ore of 3：1,temperature of 1200℃,tablet-forming pressure of6 MPa,and roasting time of 2 h.The decomposition of CaSO4reduced the potassium extraction rate;therefore,the required amount of CaSO4was more than the theoretical amount.However,excess CaSO4was also undesirable for the potassium extraction reaction because a massive amount of SO2and O2gas were derived from the decomposition of CaSO4,which provided poor contact between the reactants.The SO2released from CaSO4decomposition can be effectively recycled.

Chemical Evolution of LiCoO2 and NaHSO4·H2O Mixtures with Different Mixing Ratios During Roasting Process

Mixtures of NaHSO4·H2O and LiCoO2 extracted from spent lithium-ion batteries were prepared with molar ratios of 1：1, 1：2 and 1：3. The chemical evolution of the LiCoO2 and NaHSO4-H20 mixtures during the roasting process was investigated by means of thermogravimetric analysis and differential scanning calorimetry （TG-DSC）, X-ray diffraction（XRD）, scanning electron （XPS）. The results show that the chemical reactions in microscopy（SEM）, and X-ray photoelectron spectroscopy the LiCoO2 and NaHSO4·H2O mixtures proceed during the roasting process. The Li element in the product of the roasting process is in the form of LiNa（SO4）. With the increase of the proportion of NaHSO4·H2O in the mixtures, the Co element evolves as follows： LiCoO2→Co3O4→Na6Co（SOa）4→Na2Co（SO4）2. The roasting products exhibit dense structures and irregular shapes, and the bonding energy of Co increases.

Selective lithium recovery from black powder of spent lithiumion batteries via sulfation reaction:phase conversion and impurities influence

The aim of this study is to present a new understanding for the selective lithium recovery from spent lithium-ion batteries(LIBs)via sulfation roasting.The composition of roasting products and reaction behavior of impurity elements were analyzed through thermodynamic calculations.Then,the effects of sulfuric acid dosage,roasting temperature,roasting time,and impurity elements were assessed on the leaching efficiency of valuable metals.Characterization methods such as X-ray diffraction(XRD),scanning electron microscopy-energy dispersive spectroscopy(SEM-EDS),and X-ray photoelectron spectroscopy(XPS)were employed to analyze the phase transformation mechanism during roasting process.The results indicate that after sulfation roasting(n(H_(2)SO_(4)):n(Li)=0.5,550℃,2 h),94%lithium can be selectively recovered by water leaching and more than 95%Ni,Co,and Mn can be leached through acid leaching without the addition of reduction agent.During the sulfation roasting process,the lithium in LiNi_(x)Mn_(y)Co_zO_(2)is mainly converted to Li_(2)SO_(4),while the Ni,Co and Mn are first transformed to sulfate and then converted into oxide form.In addition,impurity elements such as Al and F will combine with lithium to form LiF and LiAlO_(2),which will reduce the leaching rate of lithium.These results provide a new understanding on the mechanisms of phase conversion during sulfation roasting and reveal the influence of impurity elements for the lithium recovery from spent LIBs.