Acid Leach Residue Research Articles

Separation of arsenic and extraction of zinc and copper from high-arsenic copper smelting dusts by alkali leaching followed by sulfuric acid leaching

Separation of arsenic and extraction of zinc and copper from high-arsenic copper smelting dust (HACSD) is one of the difficulties in the world. A two-stage countercurrent alkali leaching followed by one-stage sulfuric acid leaching for arsenic separation and zinc and copper extraction was developed in this work. The dusts properties, metals leaching behaviors and their phase transformations in the alkali leaching and sulfuric acid leaching were explored systematically. Results indicate that >95.5% of As was successfully separated from HACSD into alkali leachate. Arsenic and metals (Zn, Pb, Bi, Sn) contents were up to 61.50 g·L−1 and as low as 0.001–0.21 g·L−1 in alkali leachate, respectively. >97.2% of Zn and >96.4% of Cu were extracted from the alkali leaching residues into sulfuric acid leachate. The sulfuric acid leaching residues were enriched with 34.16% of Pb, 8.38% of Bi and 11.54% of Sn in oxides and sulfates phases. The metals leaching behaviors, phase transformations and microstructure of metal compounds during the leaching processes were confirmed by X-ray diffraction and scanning electronic microscopy/energy dispersive spectroscopy. The proposed process offered high separation efficiency of arsenic, zinc and copper, as well as low environmental pollution.

Biocrude Production via Non-Catalytic Supercritical Hydrothermal Liquefaction of Fucus vesiculosus Seaweed Processing Residues

The potential of using cold water brown macroalgae Fucus vesiculosus for biocrude production via non-catalytic supercritical hydrothermal liquefaction (HTL) was studied. Demineralization, residue neutralization, and high value-added product (alginate and fucoidan) extraction processes were carried out before using the biomass for HTL biocrude production. Acid leaching was carried out using three demineralization agents: distilled water, dilute citric acid solution, and the diluted acidic aqueous by-product from a continuous HTL pilot facility. Alginate was extracted via H2SO4 and NaCO3 bathing, and fucoidan was extracted using CaCl2. Experimental data show that none of the leaching agents was greatly efficient in removing inorganics, with citric acid leaching with extensive neutralization reaching the highest ash removal efficiency of 47%. The produced 6 sets of biocrudes were characterized by elemental and thermogravimetric analyses. Short (10-min retention) HTL and the extent of leaching residue neutralization were also investigated. Highest biocrude yields were recorded when liquefying non-neutralized citric acid leaching, alginate, and fucoidan extraction residues. On the other hand, thermochemical conversions of short retention time HTL, full neutralization extent, and baseline (dried raw macroalgae) biomass performed worse. Specifically, the highest biocrude yield of 28.2 ± 2.5 wt.% on dry ash-free feedstock basis was recorded when liquefying alginate extraction residues. Moreover, the highest energy recovery of 52.8% was recorded when converting fucoidan extraction residues.

Recovery of Zinc and Silver from Zinc Acid-Leaching Residues with Reduction of Their Environmental Impact Using a Novel Water Leaching-Flotation Process

Zinc-leaching residue (ZLR) is a strongly acidic hazardous waste; it has poor stability, high heavy metal levels, and releases toxic elements into the environment. ZLR has potential as a valuable resource, because it contains elevated levels of zinc and silver. In this paper, the recovery of zinc (Zn) and silver (Ag) from ZLR wastes from zinc hydrometallurgy workshops using water leaching followed by flotation was studied. During water leaching experiments, the zinc and copper recovery rates were 38% and 61%, respectively. Thereafter, various flotation testing parameters were optimized and included grinding time, reagent dosages, pulp density, flotation time, and type of adjuster. Experimental results demonstrated this flotation method successfully recycled Ag and Zn. A froth product containing more than 9256.41 g/t Ag and 12.26% Zn was produced from the ZLR with approximately 80.32% Ag and 42.88% Zn recoveries. The toxicity characteristic leaching procedure (TCLP) results indicated the water-leaching flotation process not only recycled valuable metals such as zinc and silver in zinc-containing hazardous wastes but lowered the hazardous waste levels to those of general wastes and recycled wastes in an efficient, economical, and environmentally friendly way.

Reaction mechanism of antibiotic bacteria residues as a green reductant for highly efficient recycling of spent lithium-ion batteries

In this study, antibiotic bacteria residues (ABRs) is proposed as a novel green reducing agent and it is found that the reducing agent can realize the efficient recovery of the valuable metals in spent cathode powder (SCP), reduce the environmental pollution and realize the high-value utilization of the biomass waste. The leaching efficiency of Ni, Co, Mn and Li can reach 99.57%, 98.50%, 98.99% and 99.90% respectively under the optimal conditions of 3 mol L−1 H2SO4, mass ratio of ABRs to SCP of 0.8:1, liquid/solid ratio of 30:1 mL g−1, the temperature of 363 K and time of 2.5 h. Leaching kinetics results shows that the reaction process is controlled by the chemical reaction with apparent activation energy exceeding than 40 kJ/mol. More importantly, the detailed ABRs leaching mechanism is proposed that the metabolite of CaC2O4 and reducing sugar in ABRs provide a synergistic reduction effect on the recovery of valuable metals. Furthermore, acid leaching residue can be regenerated to obtain lithium-ion anode materials with excellent electrochemical properties. The entire process is a sustainable green recycling strategy by using waste ABRs waste to treat spent lithium-ion batteries (LIBs), recovering valuable metals efficiently and minimizing environmental pollution.

Extraction of Sc from Sc-Bearing V–Ti Magnetite Tailings

In this study, a magnetic separation–shaking table gravity separation process was used to treat Sc-bearing V–Ti magnetite tailings, and the Sc2O3 content in the Sc concentrate was enhanced from 0.0039 to 0.0061%. The majority of the Sc in the Sc concentrate originates from augite, hornblende, and aluminosilicate. Magnesium chloride and the Sc concentrate were placed in a roasting furnace for chloridizing roasting, causing the partial chlorination of Sc to ScCl3. Sc was extracted from the roasted ore using hydrochloric acid leaching. Sc extraction indices demonstrated 95.12% Sc2O3 leaching efficiency and 0.0011% Sc2O3 content in the acid leaching residue, clearly indicating the extraction of Sc. No noticeable peak of Sc was observed in the spectrum of the acid leaching residue. Most of the Sc in the roasted ores was dissolved as Sc3+, and it entered the liquid phase. The main minerals in the acid leaching residue were Mg2SiO4, CaTiO3, and Fe2SiO4.

An innovative approach to recycle boron waste by mesoporous silica production and its application in methylene blue removal

The study shows an innovative method for boron waste treatment through synthesizing mesoporous silica (MCM-41) from it. At the initial stage of the process, magnesium was extracted from the waste by HCl solution and the silica source was leached by sodium hydroxide solution from the residue. Afterwards, n-hexadecyltrimethyl ammonium bromide (CTAB) was employed as the template. XRF investigated the influence of HCl/boron waste ratio, reaction time and temperature on preparing silica source. XRD results showed the compositions of boron waste and the acid leaching residue. Four typical characteristic peaks of MCM-41 were also appeared in the XRD results. The rod shape and 2D hexagonal pore structure of synthesized MCM-41 were determined by SEM and TEM. And it possessed a high BET surface area of 1035.534 m2/g and a large average pore diameter of 2.969 nm according to the N2 physisorption results. FT-IR result indicated calcination is an effective way to remove the template. These properties suggested synthesized MCM-41 has a wide range of applications. The adsorption capacity for methylene blue (MB) on BW-MCM-41 was explored by adsorption kinetics and isotherm models fitting as well as thermodynamic study. The effect of MB solution’s pH and the adsorbent dose was studied to determine the optimal adsorption conditions.

Phosphorus recovery from incinerated sewage sludge ash (ISSA) and reutilization of residues for sludge pretreated by different conditioners

• Effect of different sludge conditioners on P recovery from ISSA was first studied. • Fe and Ca elements in sludge conditioner influenced the P speciation in ISSA. • Higher Ca content in ISSA resulted in lower P leaching efficiency. • Heavy metals mainly existed in residual specie with low mobility risk in ISSA. • Acid-leaching residue of ISSA could be reused as cement mortar additives. Phosphorus (P) recovery from incinerated sewage sludge ash (ISSA) has attracted increasing attentions in recent years. In this study, the P recovery was comparatively investigated on two types of ISSA for sewage sludge pretreated by Fenton's reagent, and a composite conditioner of Fenton's reagent and lime (referred to as ISSA-F and ISSA-FL, respectively). ISSA derived from raw sludge (referred to as ISSA-RS) was studied as the control group. The P species results showed that the proportions of AP (Ca-bound P) and NAIP (Al/Fe-bound P) in the ISSA samples were significantly affected by the added Fe and Ca elements in the sludge conditioner. The order of leaching efficiency of P in different ISSA samples was ISSA-F (100%) > ISSA-RS (94.4%) > ISSA-FL (83.1%), which was significantly related to the Ca content in those ISSA samples. Higher Ca content in the ISSA-RS and ISS-FL samples induced calcium oxalate precipitation during leaching process, which hindered the release of P from ISSA. P recovery product mainly composed of aluminum hydroxyphosphate was prepared from the purified solutions leached from different ISSA samples. Moreover, acid-leaching residues of different ISSA samples after P recovery were proposed to be reutilized as cement mortar additives by replacing 15 wt% of cement. The mechanical strength of the test cement mortar specimens with acid-leaching residues was comparable to that of the control cement mortar specimen without residue.

Factors Affecting Alkali Activation of Laterite Acid Leaching Residues

In this experimental study, the alkali activation of acid leaching residues using a mixture of sodium hydroxide (NaOH) and alkaline sodium silicate solution (Na2SiO3) as activators is investigated. The residues were also calcined at 800 and 1000 °C for 2 h or mixed with metakaolin (MK) in order to increase their reactivity. The effect of several parameters, namely the H2O/Na2O and SiO2/Na2O ratios present in the activating solution, the pre–curing time (4–24 h), the curing temperature (40–80 °C), the curing time (24 or 48 h), and the ageing period (7–28 days) on the properties of the produced alkali activated materials (AAMs), including compressive strength, porosity, water absorption, and density, was explored. Analytical techniques, namely X-ray diffraction (XRD), Fourier transform infrared (FTIR) spectroscopy, scanning electron microscopy (SEM), and elemental mapping analysis were used for the identification of the morphology and structure of the final products. The experimental results show that the laterite acid leaching residues cannot be alkali activated in an unaltered state, and the compressive strength of the produced AAMs barely reaches 1.4 MPa, while the mixing of the residues with 10 wt% metakaolin results in noticeably higher compressive strength (41 MPa). Moreover, the calcination of residues at 800 and 1000 °C has practically no beneficial effect on alkali activation. Alkali activated materials produced under the optimum synthesis conditions were subjected to high temperature firing for 2 h and immersed in distilled water or acidic solution (1 mol L−1 HCl) for 7 and 30 days in order to assess their structural integrity under different environmental conditions. This study explores the potential of alkali activation of laterite leaching residues amended with the addition of metakaolin for the production of AAMS that can be used as binders or in several construction applications in order to enable their valorization and also improve the environmental sustainability of the metallurgical sector.

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 °C, 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.

Developed Process Circuit Flowsheet of Al Amar Ore for Production of Nanocrystalline Ferrite and Improving Gold Recovery.

Al Amar gold ore is rich in sulfides of base metals and is commercially applied for the production of copper concentrate via floatation and gold bullion by cyanidation of tailing. The current process flowsheet suffers from low gold recovery (∼60%) and loss of metals in the hazardous stockpiled residue. This work addresses these drawbacks by a newly experimental redesign of the process circuit. The innovative flowsheet comprises a sequence of operations, including acid leaching of the roasted ore, gold recovery from the leach residue, and preparation of a valuable zinc–copper–lead ferrite from the filtrate by coprecipitation followed by heat treatment. The ore is roasted at 650 °C and then leached in 20% HCl, where most of Zn, Cu, Pb, and Fe contents are dissolved, while pristine gold remains in the residue. Most of the gold (∼93%) can be recovered by cyanidation of the acid leach residue. Stoichiometric ratios of dissolved Zn, Cu, Pb, and Fe in the acid leach solution can be kept at 0.6:0.3:0.1:2.0, respectively, only by adding a small amount of ferric chloride. These metals are coprecipitated at varying pH values from 8 to 10, and the produced powders are annealed at temperatures from 600 to 1100 °C. X-ray diffraction (XRD) charts reveal sharp peaks of the targeted Zn0.6Cu0.3Pb0.1Fe2O4 phase at 600 °C, while a highly crystalline single phase is obtained at 1100 °C, independently of precipitation pH. The crystalline size of the produced powders increases with annealing temperatures (from 18–27 nm at 600 °C to 85–105 nm at 1100 °C). The finest size is found at pH 12. Scanning electron microscopy (SEM) investigation shows uniform cubic microstructures of samples annealed at 1100 °C. The produced ferrite powders exhibit soft magnetic characteristics. Saturation magnetization, Ms, substantially increases with pH. Coercivity, Hc, increases with increasing annealing temperatures, from 600 to 800 °C, and decreases above 800 °C. Preliminary cost–benefit analysis revealed that the profit margin of the proposed process flowsheet is promising. The wastewater is almost free of heavy metals. Our advances in high gold recovery and preparation of valuable magnetic nanocrystalline ferrite provide exciting opportunities to enhance and maximize Al Amar ore production for practical applications.

Optimization on the purification of cassiterite from low-grade cassiterite concentrate

Optimization on the purification of cassiterite from low-grade cassiterite concentrate has been successfully done. The concentrate used came from the mine site in the Delas area, Bangka Island with an element content of Sn was 31.60% wt and the impurity phases including zircon, ilmenite, quartz, hematite, and monazite. The purification process consists of two steps: mechanical-magnetical separation and acid leaching. The mechanical-magnetical separation step started by color separation (bright sand removal) and removed the magnetic elements of the concentrate subsequently. The content of the Sn element in acid leaching residue increased significantly was up to 90.28% with cassiterite as the dominant phase. The X-ray diffraction (XRD) revealed that the impurity phase of the filtrate after crystallization in all varying pH consists of cassiterite, ilmenite, and ammonium chloride. The solubility of cassiterite concentrate in the acid leaching process is very low at 4.4%, this indicates that the Sn element is difficult to decompose under these conditions.

Efficient separation of vanadium, titanium, and iron from vanadium-bearing titanomagnetite by pressurized pyrolysis of ammonium chloride-acid leaching-solvent extraction process

A novel process featuring pressurized pyrolysis of ammonium chloride-acid leaching-solvent extraction was carried out to separate vanadium, titanium, and iron from vanadium-bearing titanomagnetite (VBT). Then vanadium and iron existed in the form of ammonium polyvanadate ((NH4)2V6O16) and ammonium jarosite (NH4Fe3(SO4)2(OH)6) after precipitation from the water phase and the raffinate, while titanium remained in the acid leaching tailings. When pyrolysis temperature of 320 °C, pyrolysis time of 2 h, and the mass ratio of NH4Cl to slag of 1.5 followed by leaching in 61.7 g/L HCl solution, the extraction rates of vanadium, titanium, and iron reached 95.1%, 1.2%, and 92.5%, respectively. The analysis through XRD and SEM-EDS proves that the main phases of the roasted sample were attributed to FeCl2 and FeCl3, indicating pressurized pyrolysis of ammonium chloride could promote the capture of more vanadium by generated ferric/ferrous chloride in situ. The concentration of vanadium and iron was enriched to 17.8 g/L and 10.6 g/L through the N235 solvent extraction-Na2SO4 washing-Na2CO3 stripping processes, which could produce V2O5 with a purity of 99.93% by roasting (NH4)2V6O16 and ammonium jarosite with excellent crystal faces. Additionally, the TiO2 concentration reached 184.2 g/L using 80 wt% sulfuric acid for the acid leaching residue, which could be used as the raw material to production of TiO2 pigment. Moreover, the mechanism of the reaction of NH4Cl with valuable metal in the closed environment has been systematically studied.

Recovering Scandium from Scandium Rough Concentrate Using Roasting-Hydrolysis-Leaching Process

In this study, a roasting-hydrolysis-acid leaching process is used to extract scandium from the scandium rough concentrate. The scandium rough concentrate containing Sc2O3 of 76.98 g/t was obtained by magnetic separation, gravity separation, and electric separation from Sc-bearing Vi-Ti magnetite tailings in the Panxi area of China. The majority of scandium in scandium rough concentrate mainly occurs in diopside, titanopyroxene, montmorillonite, chlorite, talc, aluminosilicate minerals, and isomorphism. Sodium salt and scandium coarse concentrate are added into the roasting furnace for roasting, which makes the fusion reaction of silicon, aluminum and sodium salt to produce soluble salts such as sodium silicate and sodium metaaluminate. Scandium is further recovered from the hydrolysis residue by acid leaching. Test results show scandium leaching recovery of 95.12% and the acid leaching residue with Sc2O3 content of 8.12 g/t are obtained, while the extraction of scandium is obvious. There is no obvious peak value of Scandium spectrum in hydrochloric acid leach residue. Most of scandium in hydrolytic residue is dissolved into Sc3+ and enters into the liquid phase. The main minerals in leach residue are perovskite, ferric silicate, and olivine.

Hydrometallurgical Investigation of Sericite Clay for Extraction of Alumina and Potash Under Controlled Conditions

The current study investigates sericite clay as an alternative source of potash and alumina. The alumina and potash content in the sericite is approximately 35 and 10%, respectively, compared with 50–55% in bauxite ore. Muscovite and illite phases are stable, and direct leaching is found to be ineffective. Here, planetary ball milling was performed, which provided dissociation of stable phases to relatively reactive phases. The milling led to a five-fold increase in the surface area, resulting in enhanced leaching kinetics even at low acid concentration. The statistical design obtained optimal conditions of 8 h of milling followed by leaching in 1.73 M HCl at 75 °C, yielding 82% alumina extraction in precipitate with 85% purity in γ-alumina form and 63% potassium dissolution in sylvinite phase. The yield of the precipitate was ~65–68%, and the acid leach residue contained primarily silica.

Effective Separation and Beneficiation of Iron and Chromium from Laterite Sulfuric Acid Leach Residue

The utilization of laterite sulfuric acid leach residue has always been a challenge in the metallurgical industry. The current treatment involves deep-sea landfills, which leads to considerable environmental pollution and resource wastage. In this work, an effective method is proposed to recover iron and chromium from sulfuric acid leach residue. The effects of roasting temperature, reducing agent dosage, and reaction duration on the recovery of iron and chromium were investigated. The optimum conditions were as follows: roasting temperature of 1200 °C, reducing agent dosage of 30%, and reaction duration of 60 min. Results show that the concentrate contains 87% Fe and that only 0.7% S can be obtained with an iron recovery of 82%, indicating suitability for ironmaking. Meanwhile, 85% of chromium can be beneficiated in tailings at a chromium content of 3.5%. The microstructure and composition of reduction products were analyzed by XRD, SEM, and EDS to understand the phase transformation and the aggregation of the reduction products. Hematite in the leach residue is first reduced to FeS and Fe3O4 at low temperature and then reduced to metallic iron with increased temperature. Sulfur is combined with alunite to form FeS at a low temperature, and the decomposition of chromite at high temperature leads to the formation of chromium oxide. In a high temperature and reducing atmosphere, Cr2O3 reacts with FeS to transform Fe and CrS. Through the reduction roasting–magnetic separation process, most of the iron is beneficiated in the concentrate, and nonmagnetic minerals such as FeS and CrS remain in the tailings. Consequently, the separation and beneficiation of iron and chromium occur. The process is economically feasible and environmentally friendly and thus has great potential for the industrial-scale recycling of laterite sulfuric acid leach residue.

A novel process for recovery of aluminum, iron, vanadium, scandium, titanium and silicon from red mud

A novel process for recovery of aluminum, iron, vanadium, scandium, titanium and silicon from red mud was proposed. The parameters of HCl leaching, solvent extraction, polymerization process, alkali leaching with pressure, and aging process were investigated. The results show that more than 96 % of scandium and vanadium could be leached by using HCl with many aluminum, iron and few titanium and silicon from red mud. More than 97 % of vanadium and scandium was extracted by solvent extraction with mono(2-cthylhexyl)2-cthylhexyl phosphonate (P507), but few other elements from the acid leaching solution. The rich vanadium liquid and rich scandium liquid could be obtained by washing the saturated organic phase. More than 97 % of iron and aluminum could be precipitated as the polyaluminum ferric chloride from the raffinate. The rich titanium material of TiO2 with grade of 62 % was obtained by NaOH leaching with pressure from the acid leaching residue. The white carbon black with purity of 99.5 % was obtained by aging with sulfuric acid solution.

Selective removal of cobalt and copper from Fe (III)-enriched high-pressure acid leach residue using the hybrid bioleaching technique

Removal of metals from high pressure acid leaching (HPAL) residue was essential to alleviate potential environmental threat and avoid valuable metals loss. However, cost-effective metals extraction from HPAL residue remains a difficulty. In this study, a hybrid bioleaching process was developed for Co and Cu extraction from HPAL residue of Cu-Co sulfide ores. Results for microbial community structure optimization showed that moderate thermophilum consortium with coexistence of iron oxidizer and sulfur oxidizer was more efficient on metal extraction compared with mesophiles. Further addition of citric acid, Fe (II) and S0 significantly enhanced the release of metals through improving the total biomass, attached cells and community diversity. As a result, 87.91% of cobalt and 58.52% of copper were extracted at initial pH 1.4 and pulp density of 50 g/L by hybrid bioleaching. The hazardous potential assessments revealed that the bioleached residue could be disposed safely. These findings demonstrated that organic acids assisting bioleaching with community adjusting was a promising strategy for metals removal from HPAL residue.

Ti leaching differences during acid leaching of coal gangue based on different thermal fields

Two thermal fields, namely, microwave assistance and traditional heating, were used in this study to extract valuable Ti mineral resources from industrial solid waste (i.e., coal gangue) through acid leaching. Moreover, leaching extraction differences between the two thermal fields were analyzed. The leaching optimization experiment in different thermal fields revealed that the microwave power, acid-to-gangue ratio, and microwave radiation time under the microwave thermal field were 800 W, 1.5, and 1.25 h, respectively; the Ti leaching rate in coal gangue under these conditions reached 81.25%. By contrast, the heating temperature, acid-to-gangue ratio, and heating time under the traditional thermal field were 170 °C, 1.4, and 4 h, respectively; the Ti leaching rate in coal gangue under these conditions reached 91.83%. Characterization analysis demonstrated that the phase compositions of Ti in acid-leached coal gangue under the microwave thermal field were Ti2H2(SO4)4·5H2O and H2Ti3O7, whereas those under the traditional thermal field were H2TiO3 and Ti2(SO4)3. Ti-containing compounds were detected from acid leaching residue under the microwave thermal field, whereas those under the traditional thermal field were mainly silicon oxide and calcium sulfate. The SEM graph of acid leaching residue also reflected microstructural differences in acid leaching residue under the microwave and traditional thermal fields. This study provides new perspectives in the highly efficient and comprehensive utilization of Ti mineral resources in coal gangue.

The geologic history of Vesta inferred from combined 207Pb/206Pb and 40Ar/39Ar chronology of basaltic eucrites

Eucrites represent samples of mafic crust of the parent body, likely Vesta, and record a complex geologic history involving magmatism, metamorphism, impact processing, and metasomatism. For better understanding of the geologic history, we present combined 207Pb/206Pb and 40Ar/39Ar chronology of four basaltic eucrites showing distinct petrologic features: Agoult, an equilibrated eucrite with granulitic textures; Camel Donga, a brecciated equilibrated eucrite enriched in metallic iron; DaG 380, a partially equilibrated monomict eucrite; and NWA 049, a metasomatized polymict eucrite comprising mostly unequilibrated clasts. The 207Pb/206Pb dating in combination with the acid washing technique was applied to pyroxene-rich and plagioclase-rich fractions, while the 40Ar/39Ar dating was performed mainly on plagioclase grains using the laser incremental heating method. The Pb isotopic data for acid leaching residues are more radiogenic than those for acid washes, reflecting efficient removal of non-radiogenic Pb components by the acid washing. Among the residues, however, only those of the Agoult and Camel Donga plagioclase-rich fractions yielded isochrons, rather than errorchrons, with 207Pb/206Pb ages of 4532.39 ± 0.87 Ma and 4515.43 ± 0.42 Ma, respectively. The trace element systematics further suggest that these residues essentially represent contaminant oxide minerals such as ilmenite and chromite rather than plagioclase. On the other hand, the Ar isotopic data for Agoult and Camel Donga yielded plateau ages with weighted means of 4494 ± 9 Ma and 3749 ± 25 Ma, in contrast to those for DaG 380 and NWA 049 that did not return any plateau. The 207Pb/206Pb and 40Ar/39Ar ages of Agoult are interpreted to reflect the time of the U–Pb and K–Ar system closure during a prolonged thermal metamorphism. By combining the ages reported here and in the literature with the calculated closure temperatures, the cooling rate of Agoult is determined to be 11 ± 2 °C/Ma. This cooling rate corresponds to a burial depth of 10–20 km, indicating that granulitic eucrites were originally crystallized near the surface and were subsequently buried and metamorphosed. The ∼20 Ma younger 207Pb/206Pb ages of Camel Donga than Agoult are considered to reflect delayed burial metamorphism in the source region, whereas the plateau 40Ar/39Ar age would record the time of an impact causing brecciation and efficient Ar-degassing of the buried crust. Furthermore, the model 207Pb*/206Pb* dates calculated for acid-washed pyroxenes imply that the burial metamorphism of Agoult and DaG 380 were broadly contemporaneous and also that NWA 049 was metasomatized by fluids within the uppermost crust later than 4.4 Ga.

Novel extraction of valuable metals from circulating fluidized bed-derived high-alumina fly ash by acid–alkali–based alternate method

The recycling of valuable metals from circulating fluidized bed (CFB)-derived high-alumina fly ash (HAFA) paves the way for ash utilization with high economic benefits. However, obtaining an ideal extraction ratio for metals using a single acid treatment under moderate conditions is difficult. A moderate acid–alkali–based alternate method was used in this study to extract aluminum (Al) from CFB-derived HAFA. The dissolution behavior of lithium (Li), gallium (Ga), and rare earth elements (REY) was investigated. The characteristics of leaching residues were studied by XRD, 27Al and 29Si MAS NMR, BET, and SEM-EDS. Results show that Al in the units with low degree of polymerization distributed on the particle surface was preferentially dissolved in the HCl solution. HCl could not penetrate the particle interior and thus greatly limited the improvement of the Al extraction ratio. NaOH solution effectively removed SiO2 that accumulated on the particle surface and broke the Si–O–Al units in the residues. The exposed and liberated Al was further extracted with the HCl solution. Finally, 78% of Al2O3, 80% of Li, 72% of Ga, and 55% of REY in the ash were extracted with the HCl solution. Moreover, 63% of SiO2 in the ash was dissolved in the NaOH solution. The acid leaching residue presented abundant mesopores and a high specific surface area of 205 m2/g. The findings imply that the proposed method that combines the extraction of valuable metals and the preparation of mesoporous materials provides a new way to utilize CFB-derived HAFA.