The ocean poly-metallic nodule was leached by using slurry electrolysis process in HCl-NaCl medium. The leaching rates of Mn, Co, Cu and Ni in the ocean poly-metallic nodule are all above 97%. Meanwhile, the high purity of electrolytic MnO₂ is also obtained as an anode product. The effects of electrolysis electric quantity, acidity, temperature, slurry density, grain size and iron ions concentration were studied. The results show that the ocean poly-metallic nodule can be treated economically in the slurry electrolysis process.

A novel process was proposed for recovering nickel from nickeliferous laterite ores. First of all, silicon and magnesium were removed from lateritic ore by high concentration sodium hydroxide leaching and carbonation respectively, so as to enrich nickel. Then the method of ammonia carbonate leaching was adopted to recover nickel from the carbonized slag, and the remaining residue was used as a raw material for recovering iron. The effects of temperature, ammonia carbonate concentration, liquid/solid ratio and stirring speed on the recovery of nickel were examined. When the leached residue reacted with ammonia carbonate (6 mol·L⁻¹ ) in a ratio of liquid-to-solid of 5׃1 at 60 ℃ for 150 min at the stirring speed of 300 r·min⁻¹, approximate more than 95% nickel was recovered. During the whole process, there was no contamination produced and the chemical raw materials were recycled, thus the process was a green technology that having good social benefit.

A process of the selective leaching of nickel from low-sulfur Ni-Cu matte at atmospheric pressure was described. This matte was obtained from high grade Ni-Cu matte by magnetic separation, which mainly contained Ni-Cu alloy and a small quantity of sulfides. Firstly, the acid-oxygen (CuSO₄-H₂SO₄-O₂) leaching of the matte was conducted at atmospheric pressure. When the solution pH value reached 1.0−2.0, the oxygen flow was ceased. Then, the aqueous copper was rejected by cementation reaction with Ni in the alloy. The mineralogical characteristics of the matte in the process were analyzed by X-ray diffractometry, optical microscopy and scanning electron microscopy. And the effects of variations in temperature, particle size distribution, oxygen flow rate, pulp density, initial acid concentration and initial concentration of copper ion were investigated.

The application of microwave technique in the hydrometallurgy of nickel laterite ores was described. The mixture of nickel laterite ores and sulfuric acid was pre-treated by microwave irradiation. The dissolving of nickel was conducted in hot water at the atmospheric pressure. The effects of factors, such as microwave power, microwave irradiation time, and sulfuric acid dosage, were investigated. In microwave field, the migration of ionic species and/or rotation of dipolar species promote the liquid−solid reaction process due to the increased contact area of reactants and leaching reaction rate constant. Thanks to the strengthening action of microwave, the microwave-assisted leaching process has its advantages, such as higher extracting rate than conventional atmospheric leaching, and no need for high-pressure operation as high pressure acid leaching (HPAL). The ferric iron in leaching solution could be effectively removed by sodium jarosite process with a little loss of nickel in the jarosite precipitate.

The chloridizing segregation and magnetic separation of low-grade nickel laterites from Yunnan province of China was investigated. The nickel laterites were characterized by microscopic investigations, using X-ray diffractometry (XRD) and energy dispersive spectrometry (EDS) techniques. The pellets, which were prepared with magnesium chloride and coke as chloride agent and reductant respectively, were heated to a high temperature, and the pellets after cooling were crushed for magnetic separation. A series of experiments were conducted to examine the effect of chlorinating agent dosage, reductant dosage, chloridizing temperature and chloridizing time on enrichment grade of Ni and Co. The results indicate that the four factors have s, ignificant effects on the extractions of Ni and Co. The optimum conditions are as follows: the amounts of magnesium chloride and coke are 6% and 2%, respectively, chloridizing temperature is 1 253 K, and chloridizing time is 90 min. Under the conditions, extractions of Ni and Co reach 91.5% and 82.3%, respectively.

The leaching of magnesium from desiliconization slag of nickel laterite ores by carbonation process was studied. The influence of various parameters was investigated to optimize the conditions and determine the kinetics of the reaction. The results show that with increasing stirring speed, liquid-to-solid ratio and reaction time, and decreasing temperature, the leaching rate of magnesium enhances. The leaching process of the desiliconization slag in the range of 288−298 K is controlled by the surface chemical reaction model. The apparent activation energy is −20.45 kJ/mol, and the kinetics model is obtained.

The present study deals with the nickel extraction and iron dissolution from nickeliferous laterite by a process of sulfation-roasting-leaching. To optimize the roasting process, response surface methodology (RSM) was utilized which employed two-level and two-factor full factorial central composite design (CCD). The factors of roasting temperature and time were studied. Experiments were carried out for fitting two non-linear regression models of nickel extraction and iron dissolution. Predicted values obtained were close to experimental values, indicating the suitability of the models. Three-dimensional surface plots and contour plot were helpful to predicting the results by performing only limited set of experiments. An area of nickel extraction from 75% to 78% and iron dissolution from 5% to 10% is obtained by an overlaid contour plot. The samples roasted at different temperatures and for different durations were characterized by XRD, which show a good agreement to iron dissolution analysis.

In order to eliminate the pitting and improve the surface morphology of cathode nickel, the influence of additives of boric acid and dodecyl sodium sulfate (SDS) on the process of nickel electrowinning from sulfate system was studied by cathode polarization tests and nickel electrodeposition experiments. The experimental results show that the addition of boric acid can increase the cathode polarization while SDS can decrease the cathode polarization. Both boric acid and SDS are useful to eliminate the pitting on nickel deposits and improve the morphology of surface. Good deposit morphology with rare pitting and high current efficiency is favored by adding 10 g/L boric acid and 40 mg/L SDS in the electrolyte under the condition of nickel ion concentration of 80 g/L, sodium ion concentration of 10 g/L, pH of 3, current density of 220 A/m² and temperature of 70 ℃.

The construction procedure and structure of Ni²⁺ macromolecular complexes with series of poly styrene and maleic anhydride (PSt/MA) were studied. Both the copolymers and the complexes (Ni-PSt/MA) were characterized by means of FTIR, elemental analysis, gel permeation chromatography (GPC) and atomic adsorption spectrometer (AAS). A further study using PSt/MA as chelating resin to recover Ni²⁺ was performed. The adsorption behavior for Ni²⁺ and various relating parameters of PSt/MA in the separation process were determined. The results indicate that the adsorption capability varies with pH values and the PSt/MA can recover Ni²⁺ in higher adsorption rate and higher selective coefficient from a mixed solution containing light metals such as Ca²⁺ and Mg²⁺ impurities.

Through leaching from residue directly and leaching after a roasting treatment, respectively, the experimental research on sulfuric leaching of vanadium from residue of stone coal that came from power generation was conducted. Factors which influence the leaching of vanadium such as concentration of sulfuric acid, leaching temperature, leaching time and liquid-to-solid ratio were investigated in both processes. In the process of direct leaching, to achieve a leaching rate of 74.49%, H₂SO₄ concentration of up to 5.4%, leaching temperature of 90 ℃ and leaching time of 8 h were necessary reaction factors. The results show that after a roasting treatment at the optimum condition of 950 ℃ at 1 h, 76.88% vanadium can be leached under the experimental condition of 0.45% of H₂SO₄, 30 ℃ for 1 h with a liquid-to-solid ratio of 2 mL/g. Leaching after an oxidation roasting treatment is an efficient way to leach vanadium from the residue of stone coal, which has some advantages, such as high recovery, low economic cost and less impurities in leaching solution.
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Extraction of vanadium from black shale was attempted in pressure acid leaching. The chemical components of the sample obtained from Guizhou Province of China show that it contains 3.258% V₂O₅, 52.880% SiO₂ and 16.140% Al₂O₃. Phase analyses of vanadium indicates vanadium mainly exists in the free oxide and mica. Vanadium contents in the two phases are 18% and 53%, respectively. The contents of V³⁺, V⁴⁺ and V⁵⁺ are almost equal. Under the optimum parameters of one-step leaching (reaction time of 3 h, sulfuric addition of 25%, temperature of 150 ℃, liquid to solid ratio of 1.2 mL/g, catalyst (FeSO₄) addition of 5% and size of 85% particle 0.074 mm), about 77% of vanadium is recovered. After two-step countercurrent leaching, the leach recovery of vanadium can reach above 90%. Air replacing oxygen is completely feasible. The impurity metals can dissolve into solution in different degrees.

Vanadium extraction from stone-coal was investigated by oxygen pressure acid leaching and solvent extraction. The mineralogy of the stone-coal from Tongren City of Guizhou Province, China, was investigated by various determination methods. The effects of leaching time, leaching temperature, leaching agent concentration, leaching L/S ratio, granularity of material, additive consumption were investigated based on the mineralogy. The results show that under the conditions of leaching time of 3−4 h, temperature of 150 ℃, sulfuric acid consumption of 25%−30%, ratio of liquid to solid of 1.2:1, the granularity less than 0.074 mm, additive consumption of 3%−5%, and oxygen pressure of 1.2 MPa, and the vanadium leaching rate can be more than 92% by the method of two-step pressurized acid leaching. The powdery V₂O₅ product with 99.52% in V₂O₅ content is obtained by the flowsheet of acid recovery, removing iron by reduction process, solvent extraction, precipitating vanadium with ammonium water, and pyrolysis from the stone-coal oxygen pressure acid-leaching solution. The total recovery efficiency of vanadium is above 85%, which is more than 20% higher than that obtained in the conventional process. Furthermore, the new process does not cause air pollution since no HCl or Cl₂ is released by calcination of the raw material.

An effectively new technology of extracting vanadium from stone coals by high concentration sulfuric acid was researched. The effect of the concentration of sulfuric acid, leaching temperature, leaching time and helper leaching agent on the extraction of vanadium was explored. The results show that the optimal conditions of extraction are as follows: the concentration of sulfuric acid is 6 mol/L, the ratio of liquid to solid is 3׃1; the temperature is 90 ℃; the leaching time is 3−5 h, the diameter of the ore particle is less than 180 μm, and the concentration of helper leaching agent R is 6%. Under these conditions, the extraction of vanadium can reach 95.86%.

The recovery of vanadium from a black shale from Guizhou Province of China was performed using a three-step process, which consists of a leaching step in the H₂SO₄-HF-NaClO system under atmospheric pressure, the vanadium separation from leachate by solvent extraction and stripping, followed by precipitation of ammonium vanadate and thermal decomposition. Under the optimum leaching conditions of 100 g/L sulfuric, 15 g/L hydrofluoric acid, 1.5 g/L NaClO, 6 h leaching at 90 ℃ and oxygen partial pressure of 1 200 kPa, and the liquid-to-solid ratio of 2, about 91% vanadium extraction is achieved. The vanadium extraction yields of solvent and stripping are 99.83% and 97.66% when using 10% (volume fraction) P204, 5% TBP and 85% sulfonated berosene as organic phase and 15% (mass fraction) sulphuric acid as stripping agent. After thermal decomposition the purity of powder vanadium pentoxide products is 99.18% and the overall vanadium recovery is over 81% in the whole process.

The effects of reaction time, reaction temperature, stirring speed and flowrate of CO₂ gas on desilication rate and loss rate of vanadium were studied. The results indicated that desilication rate increases with the increase of flowrate of CO₂ gas and reaction time. Reaction temperature and stirring speed have little effect on desilication rate, while influence the loss rate of vanadium significantly. Under the condition of reaction time of 2 h, reaction temperature of 95 ℃, stirring speed of 180 r/min, flowrate of CO₂ gas of 60 mL/min and aging time of 2 h, desilication rate is more than 96% and the loss rate of vanadium is about 4.24%. The residue of desilication process can be processed for silicon materials, such as high-grade hydrated silica, which commonly known as white carbon black. In addition, with this carbonation method, leaching regents NaOH can be recycled by simple treatments.

Effects of particle size of the zinc sulfide concentrate, leaching temperature, solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied. The results indicate that the additive can improve the reaction kinetics and the conversion rate. And sulfur can be successfully separated from the zinc sulfide concentrate as elemental sulfur. The reasonable experiment parameters are obtained as follows: the leaching temperature 150 ℃, oxygen partial pressure     1 MPa, additive amount 1%, solid-to-liquid ratio 1:4, leaching time 2 h, initial sulfuric acid concentration 15%, and particle size less than 44 μm. Under the optimum conditions, the leaching rate of the zinc can reach 95% and the reduction rate of the sulfur can reach 90%.

Indium was recovered from zinc oxide flue dust (ZOFD) with sulfuric acid by oxidative pressure leaching in an autoclave, and the effects of different technological conditions on indium leaching were studied. Potassium permanganate and hydrogen peroxide were used as oxidants. The atmospheric pressure leaching experiments were also carried out. The experimental results show that the leaching rate of indium can be effectively improved by oxidative pressure leaching. The optimum conditions of pressure leaching are determined as sulfuric 5.10 mol/L acid, leaching time 150 min, temperature 90 ℃, and the H₂O₂ dosage of 0.5 mL/g or 2.5% KMnO₄. The leaching rate of indium is more than 90%, which is increased by 13% compared with that of atmospheric pressure leaching process without oxidant under the optimum conditions.

Copper adsorption by orange peel, which was chemically modified with sodium hydroxide, was investigated. The adsorbent was characterized using surface area analyzer, infrared spectroscopy and scanning electron microscopy. Total negative charge and zeta potentials on the adsorbent surface were determined. Equilibrium isotherms and kinetics were obtained and the effects of solution pH value, adsorbent concentration and temperature were studied in batch experiments. Column experiments were performed to study practical applicability, and breakthrough curves were obtained. Equilibrium is well described by Langmuir and Freundlich isotherms, and kinetics is found to fit pseudo-second order type adsorption kinetics. According to Langmuir equation, the maximum adsorption capacity for Cu(Ⅱ) is 50.25 mg/g at pH value of 5.3. The results show additional chemical modification of the adsorbent by NaOH and the increased adsorption capacity.

The waste water system generated in the process of production of cuprous chloride was studied. The existing forms of copper in the system and the influence of temperature and pH on the existing forms of copper ion were analyzed and determined through calculating the coefficients of copper complex distribution. In the waste water system, the main forms of copper are CuSO₄, Cu²⁺, CuCl⁺,  and . Temperature has little influence on the distribution coefficient of Cu(Ⅱ), but has significant influence on distribution coefficient of Cu(Ⅰ). With the increase of temperature, the distribution coefficient of  increases significantly while the distribution coefficient of  decreases. The pH has nearly no influence on the distribution coefficients of various Cu(Ⅰ)-compounds, but has sizable influence on the distribution coefficients of Cu(Ⅱ)-compounds. With the increase of pH, the distribution coefficient of CuSO₄(aq) increases while the distribution coefficients of Cu²⁺ and CuCl⁺ decrease. According to these results, the anion resin of  201×7 OH⁻ and the cation resin of 732 Na were chosen to dispose the waste water solution of cuprous chloride. Finally, 97.9% copper in the waste water is recovered.

The solvent of P350 was applied to extract and separate the oxalic acid from the mother-liquor originated from the precipitation of cobalt by oxalic acid, and its extraction mechanism was deduced. Some factors, including the concentration of P350, the concentration of hydrochloric acid and the concentration of oxalic acid were investigated to determine the best distribution coefficient of the oxalic acid. In the case of phase ratio (O/A ) at 2.0, the extraction of the oxalic acid was more than 95% and its concentration in the extraction raffinate was lower than 0.004 0 mol/L after six-stage counter-current extraction. While the phase ratio (O/A) of the stripping was at 1.0, the recovery of oxalic acid attained more than 95% after ten-stage counter-current stripping.

Decomposition of sodium aluminate solution is the key step for alumina production; however, approximately only 55% alumina yield is obtained by seeded precipitation. So, it is difficult to obtain sodium aluminate solution with high caustic ratio. In this study, a method was explored to recover alumina from spent Bayer liquor by deep decomposition with methanol. A variety of conditions, including reaction temperature, reaction time, methanol amount and seed coefficient were elaborately investigated. The results showed that the appropriate conditions were 1:1 in volume ratio of methanol to spent Bayer liquor, more than 1.0 seed coefficient and in a 40 ℃ water bath for 24 h. By characterizing through XRD, the crystal products were found to be Al(OH)₃. With this method, the molar ratio of Na₂O to Al₂O₃ of the spent liquor can be increased from about 3.0 to 10.0 due to the recovery of alumina, which is beneficial for the treatment of red mud.

By desilication treatment, the Al₂O₃/SiO₂ molar ratio of coal fly ash could be improved to the range of 1.63−2.0. The desilicated coal fly ash (DSCFA) was enriched in alumina extraction. A processing technology was developed for alumina extraction from the DSCFA with the lime sintering process. Ca/(SiO₂+TiO₂) molar ratio, and NaO/Al₂O₃ molar ratio, sintering time, and temperature were the most significant parameters impacting on the aluminum extraction efficiency. The optima aluminum extraction efficiency was obtained under conditions of Ca/(SiO₂+TiO₂) molar ratio of 2.0, NaO/Al₂O₃ molar ratio of 0.98, and sintering at 1 200 ℃ for 60 min. A standard industrial dissolution method was used under conditions of caustic ratio (α_k=n(NaO)/n(Al₂O₃) of 2.0, Al₂O₃ concentration of 50 g/L, sodium hydroxide concentration(N_k) of 60.78 g/L, Na₂CO₃ concentration of 10 g/L, temperature of 85 ℃, and dissolution duration of 10 min. The final aluminum extraction efficiency was 90%.

The application of pressure leaching technology in the treatment of high-copper and high-arsenic dust was studied. The pressure leaching technique was determined as follows: the liquid to solid ratio (mL/g) of 5:1, the leaching temperature of 453 K, the retention time of 2 h, the initial sulfuric acid concentration of 0.74 mol/L, the oxygen partial pressure of 0.7 MPa, and the agitation speed of 500 r/min. Under these conditions, 95% of copper and 99% of zinc and only 6% of iron in the dust were leached, while about 20% of arsenic was also leached. The leaching technique was optimized further to restrain the leaching of arsenic by adding a small quantity of ferrous iron into the leaching system (c(Fe²⁺)=0.036 mol/L). Copper and zinc can be effectively separated from arsenic and iron in the leach. The optimal pressure leaching technique of high-copper and high-arsenic smelter dust is proved to be effective.

A new technology for microwave pretreatment of spent catalyst on Zn extraction by HCl leaching was proposed and the temperature-change curve of spent catalyst under microwave irradiation was measured. The influence of microwave pretreatment temperature and microwave irradiation time on zinc extraction was investigated and the mechanism of microwave pretreatment for spent catalyst was analyzed. The results show that microwave pretreatment can greatly enhance the leaching rate of Zn. The Zn extraction reaches 96.58% under the conditions of microwave pretreatment temperature of 950 ℃ and the microwave irradiation time of 12 min. The blocked pores of spent catalyst can be opened through microwave pretreatment, increasing the contact area of leaching reagent and zinc.

Preparation of orange peel xanthate and its adsorption behaviors of five heavy metals (Cu²⁺, Cd²⁺, Pb²⁺, Zn²⁺ and Ni²⁺) were studied. FTIR spectra, Zeta potentials and TG analysis were used to characterize prepared orange peel xanthate. Effects of various parameters including equilibrium pH, initial metal ion concentration and adsorption time on the adsorption processes for the five metal ions were investigated. It was found that for all five metal ions, the adsorption isotherms agreed Langmuir model very well and the maximum adsorption capacities of Cu²⁺, Cd²⁺, Pb²⁺, Zn²⁺ and Ni²⁺ were obtained as 77.60, 76.57, 218.34, 49.85 and 15.45 mg/g, respectively. All adsorption processes can attain equilibrium within 20 min and kinetics was well fitted by psesudo-second order equation. It is proposed that the adsorption mechanism was complexation.

Silver as a highly conductive metal is usually doped in YBa₂Cu₃O_7−x superconductor bulks to improve critical current density of YBa₂Cu₃O_7−x superconductor. The valuable metal elements silver, yttrium, barium and copper in waste YBa₂Cu₃O_7−x/Ag composite superconductor bulks were recovered, respectively. Silver was recovered with process at first, the waste was dissolved by nitric acid and silver chloride was precipitated by adding chloride acid, then silver pig was obtained by melting silver chloride together with sodium carbonate at 1 000 ℃. The effective factors on recovery ratio and purity of silver were studied. The chemical analysis proves that the purity of silver ingot is 95.86%. The recovery ratio of silver is calculated to be 92.56%. The loss of silver may be due to the loss of silver chloride during filtering and the volatilization of silver when silver chloride and sodium carbonate are smelted at high temperature. For other three metal elements, Y³⁺, Ba²⁺ and Cu²⁺, in the surplus waste liquid after recovering silver, they were separated with the sequence of barium, copper and then yttrium step by step. First, sulfate acid was used to precipitate barium sulfate. Then, sodium sulfide was added to the surplus solution so that copper could be separated as copper sulfide. During this separation procedure, it was important to control the pH value to be 1−2. After that, oxalic acid was added into the surplus solution to obtain yttrium oxalate. Finally, yttrium oxide was formed by burning yttrium oxalate. The XRD results indicate that the final products are all single-phase compounds as BaSO₄, CuS and Y₂O₃, respectively.

A greenness evaluation index and system of microwave-assisted leaching method were established. The effects of the life cycle assessment variables, such as the resource consumption, environment impact, cost, time and quality, were investigated, and the concept of green degree was applied in the production of synthetic rutile. An analytic hierarchy process was utilized to assess matrix of greenness evaluation. The Gauss-Seidel iterative matrix method was employed to solve the assessment matrix and obtain the weights and membership functions of all evaluation indexes. A fuzzy decision-making method was applied to build the greenness evaluation model, and then the scores of green degree in microwave-assisted leaching process was obtained. The greenness evaluation model was applied to the life cycle assessment of the microwave-assisted leaching process. The results show that the microwave-assisted leaching process has advantages over the conventional ones, with respect to energy-consumption, processing time and environmental protection.
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