Acid Leaching Solution Research Articles

Thermodynamics of Separating Molybdenum(VI) Over Iron(III) from High Acid Leach Solutions with Mixtures of P507 and N235

Thermodynamics of Separating Molybdenum(VI) Over Iron(III) from High Acid Leach Solutions with Mixtures of P507 and N235

Indirect in situ bioleaching is an emerging tool for accessing deeply buried metal reserves, but can the process be managed? – A case study of copper leaching at 1 km depth

Copper is a strategic raw material widely needed for electrification. One possibility to diversify the supply to answer the market demand is to produce copper with in situ technology. In this study, feasibility of in situ bioleaching of copper was tested in a deep subsurface deposit. During in situ bioleaching of copper, copper is leached using a biologically produced ferric iron solution, which is recycled back to the in situ reactor after valuable metals are recovered, after which the solution is re-oxidized by iron-oxidizing microorganisms (IOB). A rock reactor was constructed in the Rudna Mine at ca 1 km depth and the microbiology and hydrogeochemistry of the water circulated through the reactor after blasting for fracturing the rock was monitored over time. The test site was rich in carbonates requiring large quantities of acid to remove the buffering capacity. The bacterial, archaeal and fungal communities in the rock reactor were monitored and characterized by quantitative polymerase chain reaction (qPCR) and amplicon sequencing, and acidophilic, iron oxidizing activity of the microbial communities during operation and pre- and post-operation phases was tested by cultivation. No acidophilic iron oxidizers were detected in the water samples during construction of the pilot reactor. Acidic leaching solution originating from the underground ferric iron generating bioreactor (FIGB) contained acidophilic IOB, which were also viable after the leach liquor was returned from the rock reactor. In the post-operation phase, when the rock reactor was neutralized with CaCO3/Ca(HCO3)2 solution, to inhibit the acidophilic IOB, iron oxidizing microorganisms were still present in the effluent solution one week after termination of the leaching and start of neutralization. Therefore, the post-operation phase needs further attention to completely stop the activity of added microorganisms. Copper was abundantly leached during the acid wash and leaching phases, proving the concept of deep in situ bioleaching.

Organic phase modification of YW100 extraction system: Extraction of germanium using YW100 + D2EHPA + N235

In this study, germanium was selectively extracted using an organic phase containing 3% C7-9 hydroxamic acid (YW100), 15% Di(2-ethylhexyl)phosphoric acid (D2EHPA), 35% Tri(octyl-decyl) amine (N235), and 47% kerosene from the Ge-slag sulfuric acid leaching solution. Then, the solution was stripped by the 1 mol/L NH4F solution to obtain a germanium extraction ratio of 99.4% and a stripping ratio of 96.2% after three extraction and stripping stages. The total organic carbon (TOC) was used to characterize the total organic matter amount in the raffinate. Compared to the “YW100 + D2EHPA” extraction system, the “YW100 + D2EHPA + N235” extraction system reduced 80% of the water solubility loss of YW100 and improved the phase separation performance. In addition, the mechanisms of germanium extraction and phase separation improvement were studied. The coordination number of YW100 and germanium is 3, while the coordination number of D2EHPA is 1.

Study on the extraction and separation of precious metals from wastewater using a hydrophobic deep eutectic solvent

Efficient and sustainable recovery of precious metals from the acidic leaching solution of solid waste is considered to be the main solution to the scarcity of precious metal resources. In this study, a hydrophobic deep eutectic solvent (HDES) was synthesized using trioctylmethylammonium chloride (N263) and oleic acid (OA). HDES was then used for the efficient recovery of Ir, Pt, Ru, and Rh from leaching solution containing precious metals. The effects of different molar ratios between N263 and OA, the initial pH of the aqueous phase, the oscillation time, and the phase ratio O/A of the extraction process were investigated. The organic phases were characterized analytically using ultraviolet-visible spectroscopy, Fourier-transform infrared spectroscopy, and electrospray ionization–mass spectrometry to further investigate the reaction mechanisms of the extraction process. The single-stage extraction percentages of Ir(IV), Ru(III), Pt(IV), and Rh(III) ions were 97.57%, 94.26%, 99.55%, and 15.1%, respectively. The high saturation extraction capacity of HDES enabled a high precious metal load, reaching 18.44 g L−1. Hydrogen bonds were formed between Cl-, which was bound by electrostatic force in the N263 structure, and H in the -COOH group in the OA molecule. During the extraction process, metal chloride complex ions replaced Cl- and entered the organic phase as 2[C17H33-COOH···N263+]IrCl62-, 3[C17H33-COOH···N263+]RuCl63-, 2[C17H33-COOH···N263+]PtCl62-, and 3[C17H33-COOH···N263+]RhCl63-. After scrubbing for effective removal of Cu and Fe in the loaded organic phase, different systems were used for stripping. The stripping percentage of precious metal ions reached more than 96.3%, providing efficient separation of the four precious metals from wastewater.

Hydrothermal crystallization of jarosite and natrojarosite from acid leaching solution of zinc oxide dust

ABSTRACTThe separation of Zn and Fe, K, and Na ions remains a challenge in the recovery process of Zn-containing secondary resources. A hydrothermal process without adding neutralizer was developed to remove Fe, K, and Na ions from the leaching solution of zinc oxide dust. Results showed that Fe and K ions were synchronously eliminated, yielding low-contaminant jarosite products with excellent crystallinity and rosette-like morphology under the following conditions: gas pressure = 1.6 × 106 Pa, agitation speed = 600 rpm, reaction time = 3 h, and temperature ranges at 368.15 K−423.15 K. The removal efficiency of Fe and K ions reached 99.8% and 97.7%, respectively. The obtained low-contaminant jarosite contained Fe 31.09%−31.98%, Zn 1.05%−1.24%, S 12.33%−12.71%, K 6.12%−6.46%, and Na 0.44%−0.87%. Additionally, the Na ions in the leaching solution precipitated can also along with Fe ions, forming low-contaminant natrojarosite products containing Fe 31.36%−35.66%, Zn 1.60%−1.88%, S 10.66%−12.56%, Na 2.85%−3.26% and K 0.74%−0.85%. The optimal hydrothermal temperature for the removal of Fe, K, and Na ions while minimising energy consumption was determined to be 393.15 K. The low-contaminant jarosite and natrojarosite products possess potential applications in the field of yellow pigments, building materials, iron and arsenic removal, adsorption materials, catalytic materials, etc.

Removal of fluoride from the mixed Ni-Co-Mn sulfate leach solution of spent lithium ion batteries using polyaluminum sulfate

The mechanism and application of defluorination in a simulated leaching solution of spent lithium-ion batteries (LIBs) using poly‑aluminum sulfate (PAS) were studied. The experimental results indicated that reaction parameters such as the molar ratio of Al to F, pH of the solution, reaction temperature, initial concentration of fluorine, and concentration of lithium and nickel, significantly affected the reaction efficiency. Under the optimum reaction conditions, the percentage of defluorination was 95.5 (wt%). During the defluorination reaction, Al3+, Al(OH)3 reacted with F−, HF to generate AlF3 and AlF(OH)2. The crystallinity of AlF3, AlF(OH)2, and Al(OH)3 and the chemical stability of the precipitates improved with the increase in temperature and pH. The Li+ concentration in the solution increased the defluorination efficiency by forming LiAlF4. At the same time, high concentrations of NiSO4 inhibited the defluorination effect and were lost in association with the solids, NiSO4·6H2O and Ni(OH)2. According to the study of reaction kinetics, the rate equation of the aluminum‑fluorine reaction was -d[F]/dt = 0.0384exp(−8.59/RT) [F0]1.11[Al]0.31. The activation energy of the precipitation reaction was 8.59 kJ/mol, and the reaction rate was relatively high. In the actual defluorination process of the acid-leaching solution of waste battery powder with Ni-Mn-Co chemistry, the F− concentration reduced from 1205 mg/L to 48 mg/L and the loss of valuable metals was <6%, and the Ni, Co, and Mn in the precipitates were predominantly sulfates and hydroxides.

Experimental investigation on the influence of surfactant to the seepage characteristics of acid leaching solution during in-situ leaching of uranium

Experimental investigation on the influence of surfactant to the seepage characteristics of acid leaching solution during in-situ leaching of uranium

Separation of metals from acetic acid leaching solution of spent lithium-ion batteries by ionic liquid system

Spent Li-ion batteries (LIBs) pollute the environment and endanger human health; however, they are valuable mineral resources. Here, we designed an extraction system composed of 0.4 mol/L 2-thenoyltrifluoroacetone (HTTA), 0.4 mol/L trioctylphosphine oxide (TOPO), and ionic liquid (IL) 1-Butyl-3-MethylImidazolium bis (trifluoromethyl sulfonyl) imide to recover valuable metals from spent LIBs. Under the optimum conditions (O/A = 1/1, 25 °C, reaction time 20 min, and pH = 3.65, the pH of the solution does not need to be adjusted), over 99.54% of Co, 99.85% of Mn, and 99.68% of Ni were extracted, while only 2.84% of Li were coextracted. The ultraviolet spectrum analysis showed that the spectral peaks of the stripping solution changed at 266 and 293 nm, indicating that CH3COO– participated in the extraction reaction. Following stripping with sulfuric acid, the pH of the solution was adjusted with 2.0 mol/L NaOH, and approximately 99% of the Al3+ ions were precipitated. Nickel and manganese sulphates were added to the stripping solution. The precursor was prepared using a coprecipitation method. The precursor was then mixed with lithium hydroxide and calcined at 1000 °C for 12 h to obtain LiNi1/3Co1/3Mn1/3O2.

Extraction kinetics of separating molybdenum(Ⅵ) over iron(Ⅲ) from high acid leach solutions with mixtures of P507 and N235

In view of the difficulty of separating molybdenumVI over ironIII from high acid leach solutions, the extraction kinetics of separating molybdenumVI/ironIII with mixtures of 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester (P507) and long chain tertiary amine (N235) was investigated using a Lewis cell to enhance the separation efficiency. Parameters affecting the extraction process such as stirring speed, specific interfacial area, and temperature were studied. The results show that the extraction process of MoVI takes place in the bulk aqueous phase, and the extraction rate is controlled by diffusion (The apparent activation energy was calculated to be 16.66 kJ/mol). The rate constants (kf) of mean MoVI extraction was calculated as 100.3, and the extraction rate equation was established: - d[Mo]/dt = 100.3[Mo](a)[P507](o)0.35[N235](o)0.24. The extraction reaction of MoVI had a fast rate in the studied range of solution conditions, but FeIII was not extracted. Moreover, the separation mechanism was speculated by two-phase titration, Fourier transform infrared spectroscopy and Nuclearmagneticresonancespectroscopy.

A clean process for the recovery of iron as value-added product from highly acidic chloride leach solution

Recovery of potassium as potassium chloride from glauconitic rocks having generalized chemical formula of (K, Na) (Fe3+, Fe2+, Al, Mg)2 (Al, Si)4 O10 (OH)2·nH2O through reduction and redox roasting followed by HCl leaching leads to generation of highly acidic chloride solution containing K as well as Fe. The dissolved iron was present in both the form that is Fe2+ for reduction roasting and Fe3+ for redox roasting. Separation of such iron is one of the major concerns for most of the hydrometallurgical industry. Currently, the most common process for iron removal is precipitating it in the form of jarosite, para-goethite and goethite. However, this process is not suitable due to high alkali consumption, formation of metastable solid and environmental concern related to its disposal. These products don’t have any further application. In order to avoid such complication and recover it as value-added product, the authors has proposed an air oxidative precipitation method for Fe2+ and hydrothermal treatment for Fe3+ recovery from such acidic leach chloride solution. The final product of these processes are nano sized monodispersed magnetite and hematite particles. The synthesized products were characterized through X-ray diffraction (XRD), Scanning electron microscopy (SEM) followed by Energy-dispersive X-ray (EDS), Fourier transform infrared spectroscopy (FTIR) and Vibrating sample magnetometer (VSM) to confirms its suitability for pigment application.

Assessing the Efficiency of Ion Exchange Resins for the Recovery of Scandium from Sulfuric Acid Leaching Solutions

Scandium, a valuable element with restricted production sources mainly situated in China and Russia, is typically obtained as a by-product during the production of various materials. As the demand for scandium grows in the expanding aluminum and fuel cell industries, and with significant investments in rare earth mining in the USA and Australia, there is a need to explore alternative recovery sources. This research investigates the recovery of scandium from an acid pregnant leaching solution using ion exchange resins. The pregnant leaching solution was obtained after the leaching of bauxite residue with sulfuric acid. Commercial resins with different functional groups were tested for their performance in scandium extraction. In addition, the co-adsorption of impurities, such as iron and titanium, was studied. The feed solution consisted of 12.7 mg/L Sc and main impurities of 272 mg/L Fe and 33.6 mg/L Ti and was pretreated before the ion exchange process by acidification with sulfuric acid and iron powder addition to suppress silica gel formation and minimize the Fe(III) content in the solution accordingly. Among the tested resins, a D2EHPA-impregnated resin had high selectivity for Sc towards Ti, while a monophosphonic resin was also a promising option since it had a higher capacity for Sc but co-extracted Ti. These findings offer promising opportunities for the recovery of scandium from acid leaching solutions and could contribute to addressing the growing demand for this valuable element.

Study of Factors Affecting the Copper Ore Leaching Process

This paper provides an overview of hydrometallurgical copper extraction studies in which liquid extraction technology has been used with four copper deposits of different compositions. The sulfuric acid consumption rate and copper extraction efficiency, which are dependent on the initial content and forms of calcium compounds and other impurities in ore samples, were calculated, and the results are presented herein. It was established that during the leaching process, silicate compounds of alkaline earth metals, in addition to calcium and magnesium carbonate compounds, would affect the levels of sulfuric acid consumption, thereby actively lowering the acidity of the environment. Moreover, these compounds could partially sorb copper ions from sulfuric acid leaching solutions. Thus, the analysis of waste ore samples showed that residual copper is mainly contained in the form of complex silicate complexes. The presence of divalent iron compounds in the composition from one of the deposits also allowed us to perform a biochemical leaching experiment with preliminary oxidation using an Acidithiobacillus ferrooxidans bacterial culture adapted to the ore composition. The use of this biochemical method in the copper leaching process resulted in a significant reduction in sulfuric acid consumption, by 40%, and a copper recovery rate of 87.2%.

Efficient removal of iron during partial neutralization of nickel laterite acid leach solution: DFT calculation and experimental verification of mechanism

Partial neutralization is typically used to remove Fe3+ ions from the acid leach solution of laterite. However, during this process, significant amounts of Mg2+, Ni2+, Co2+, and SO2–4 are lost. With increasing initial Fe3+ concentration and solution pH, the severity of loss increases, with the precipitation of Fe3+. In this study, the mechanisms of ion co-precipitation during partial neutralization were investigated. Based on density functional theory calculations, a precipitation pathway through the formation of [Fe4(OH−)2(SO2–4)4]•[2SO2–4•yH2O]•[M2+] was proposed. The differential charge density diagram indicated that the O atoms acted as the charge acceptors and maintained the precipitate structure. The coordination between M2+ and SO2–4 could be destroyed via intense agitation, releasing approximately 100% of Mg2+, Ni2+, and Co2+ ions from the precipitates to the solution. Furthermore, the addition of cationic or anionic surfactants limited the formation of these coordination bonds. The highest retention percentages of Mg2+, Ni2+, and Co2+ in solution were 99.4, 99.6, and 99.9%, respectively, after adding 2 × 10−5 mol/L of sodium dodecylbenzenesulfonate (CTAB). In addition, a high temperature could break these coordination bonds. The retention percentages of Mg2+, Ni2+, and Co2+ increased from 82.6, 85.2, and 86.5% to 99.9, 99.6, and 99.8%, respectively, as the temperature increased from 25 to 65 °C. These results confirmed the rationality of the proposed structure of the precipitate.

A sulfonated ligand-aqueous two-phase system for selective extraction of thorium from mining wastewater: Process optimization, structural characterization and mechanism exploration

As a potential nuclear fuel, the recovery of thorium with high efficiency, high selectivity and high extraction capacity from thorium-containing wastewater is considered to be one of the effective ways to reduce pollution and solve the shortage of thorium resources. However, the problem of efficient and selective separation of thorium from the complex environment has not been completely solved. Herein, a functional ligand called arsenazo (ASA), polyethylene glycol, and sodium sulphate were used to construct an aqueous two-phase system (ATPs) for the selectivity extraction and separation of thorium from mining wastewater. Furthermore, the complex agent, pH value, temperature, equilibrium time, coexisting substances, and other conditions for the extraction and separation of thorium were optimized, indicating that 98.7 % of thorium was efficiently and precisely separated from the acid-leaching solution to the pH of 2–3 or 5–6 with 0.08 g/L of ASA within 10 min, and the interference with multiple coexisting ions and organic substances was avoided. Importantly, the analysis of water quality parameters, thermogravimetry, infrared spectroscopy, and ultraviolet–visible spectroscopy revealed the coordination relationship between thorium ions and complex agents, indicating that the extraction of thorium in the aqueous two-phase system was a cation exchange mechanism. More importantly, SEM, TEM, EDX, and XRD verified that thorium was efficiently and selectively separated by a sulfone acid ligand-aqueous two-phase system, providing quantitative information on thorium-containing product distribution of relatively high accuracy and avoiding the interference with coexisting substance. Therefore, this work provided theoretical guidance on treating and reusing thorium in wastewater and enriched the content of functional ligand-aqueous two-phase extraction systems in green production technology.

Effective recovery of Ti as anatase nanoparticles from waste red mud via a coupled leaching and boiling route.

Red mud (RM) a solid waste generated by the bauxite smelting industry, is a rich source of metal resources, especially Ti, and its recycling can bring significant environmental and economic benefits. In this study, precious metal Ti was efficiently recovered from red mud using a coupled acid leaching and boiling route for the effective separation of low-value metals. The red mud which contained mainly 10.69% Si, 12.1% Al, 15.2% Ca, 10.99% Fe, and 4.37% Ti, was recovered in five steps. First, a nitric acid solution was used to leach the metals in multiple stages, resulting in an acidic leach solution with high concentrations of Fe, Al, Ti, and Ca ions 2.7g/L, 4.7g/L, 5.43g/L, and 1.8g/L, respectively. Then, a small amount of sucrose was added as a catalyst to recover Ti from the leach solution under hydrothermal conditions, resulting in the targeted recovery of 98.6% of Ti in the form of high-purity anatase while Fe, Al, and Ca remained in the solution. Next, the Fe in solution was separated as hematite products at a temperature of 110°C and a reaction time of 4h. Similarly, the Al in the solution was separated and precipitated as boehmite by heating it at 260°C for a reaction time of 20h. Finally, the remaining Ca in solution was recovered by simple pH regulation. Economic accounting assessment showed that the method yields $101.06 for 1t of red mud treated, excluding labor costs. This study provides a novel approach to recover precious metals from metal wastes through the whole process resource recovery of solid waste red mud.

Short-Process Preparation of High-Purity V2O5 from Shale Acid Leaching Solution via Chlorination

The conventional V2O5 preparation processes include ion exchange, chemical precipitation, solvent extraction, and other processes. Given the long process and complex operation nature of traditional V2O5 production methods, we herein developed a short-process, low-temperature, and convenient operation method of isolating vanadium (in the form of V2O5) from shale acid leaching solution. The acid leaching solution was oxidized with NaClO3 and pH-adjusted with NaOH to form a vanadium-containing precipitate, which was mixed with AlCl3 (V:AlCl3 = 1:5, mol/mol) and roasted for 120 min at 170 °C to afford vanadium oxytrichloride (VOCl3) with a purity of 99.59%. In addition, the vanadium-containing precipitate was mixed with AlCl3 and NaCl (V:AlCl3:NaCl = 3:12:8, mol/mol/mol) and roasted for 120 min at 170 °C to afford VOCl3 with a purity of 99.94%. VOCl3 (purity of 99.94%) was dissolved in ultrapure water, and the solution (32 gvanadium/L) was treated with NH3·H2O (NH3:V = 1.34, mol/mol) at 50 °C for 120 min. The obtained precipitate (vanadium precipitation rate = 99.28%) was roasted at 550 °C for 3 h to afford high-purity vanadium pentoxide (V2O5) with a purity of 99.86%. Compared with the traditional hydrometallurgical method of V2O5 preparation, our method avoided solvent extraction and other undesired processes and the overall process flow is greatly shortened, thus having high practical value.

Experimental Study on Synergistic Extraction for Separating Manganese and Iron from Waste Ternary Battery Leaching Solution

In this paper, the key factors and the mechanism of the extraction of iron and manganese from the sulfuric acid leaching solution of waste ternary lithium-ion batteries were studied by combining P204 and N235 extractors. The experimental results showed that the optimal organic phase composition was 25% P204 + 15% N235 + 60% sulfonated kerosene, and the optimal pH of the pre-extraction solution was in the range of 3.0–3.5. Under the conditions of an extraction temperature of 25–35 °C and an extraction ratio of O/A = 1/1 and after mixing for 5 min, the removal rate of Mn and Fe exceeded 99%. Under the same extraction conditions, the extraction effect of the P204-N235 composite extractant on Mn and Fe was better than that of P204 alone. logD-log[P204] showed that the formulas of the Mn and Fe extracts were MnR2(HR) and FeR2(HR), respectively. The logD-pH diagram showed that only one free H+ was released for each metal ion during extraction, and extraction occurred via a cation exchange reaction.

Separation of molybdenum from peroxotungstic acid solution by solvent extraction with TBP and TRPO

An increasing problem facing the tungsten industry is developing an effective separation of molybdenum from peroxotungstic acid solutions due to the similarities of these two metals. This study herein separates molybdenum from peroxotungstic acid solution for the selective production of tungsten from scheelite acid decomposition residue using hydrogen peroxide. Solvent extraction was the main method employed, and the feed was peroxotungstic acid solution containing peroxomolybdic acid. The concentration of tungsten and molybdenum in the feed and the raffinate were measured by Inductively Coupled Plasma (ICP) to characterize the separation efficiency. The results showed that in the peroxide solution, an organic phase of tributyl phosphate (80 %), trialkyl phosphine oxide (2 %), and sulfonated kerosene (18 %) was used effectively to separate molybdenum from tungsten in acid leach solutions. A maximum single-stage separation coefficient of 400 was reached under the following extraction conditions: temperature, 25 °C; contact time 4–6 min; separation time 4 min; and organic to aqueous ratio, 1.2. Using ammonium bicarbonate as the stripping agent, the stripping ratio of molybdenum and tungsten were 83 % and 33 %, respectively, and the corresponding separation coefficient was 10.

Soil Contaminated with Hazardous Waste Materials at Rio Tinto Mine (Spain) Is a Persistent Secondary Source of Acid and Heavy Metals to the Environment

Mineralogical analysis and laboratory-based leaching tests coupled with speciation modeling were undertaken to quantify the potential for short-term acid generation and the release of trace elements from soils heavily contaminated with mine waste at Rio Tinto. Three different waste materials were considered as case studies: roasted pyrite, copper slags, and leached sulfide ores. The results showed elevated values of net acid generation (up to 663 mmol H+/kg), the major pools being potential sulfidic acidity and acidity retained in jarosite. Remarkable contents of As and toxic heavy metals were found especially in the slag-contaminated soil. Copper, Zn, and Pb were the most abundant metals in the acid leach solutions resulting from mine soil-water interaction, with peak values of 55.6 mg L−1, 2.77 mg L−1, and 2.62 mg L−1, respectively. Despite the high total contents of trace elements occurring in soil, the mobile fraction was limited to maximum release values of 12.60% for Cd and 10.27% for Cu, according to the test leaching. Speciation calculations indicated that free metal ions (M2+) and sulfate species (MSO40) accounted for most of the dissolved load. Acid soil drainage is a secondary source of acid and heavy metals in the mine site and, therefore, an effective land reclamation program should ensure that acidity and metal mobility are reduced to environmentally sustainable levels.

Investigation of a novel alkali baking–leaching process for dissolution of Fe, Al, Ti, Sc, and Ga values and generation of alumina and titania from red mud

The current study investigates alkali baking using sodium hydroxide followed by water and sulfuric acid leaching to extract Al, Fe, Ti, Sc, and Ga values. Water leaching of baked mass selectively dissolves Al. Subsequent, sulfuric acid leaching of water-leach residue dissolves Fe, Ti, Sc, Ga, and remaining Al and Si values. The alkali baking forms sodium metal oxide (NaFeO2, Na2TiO3, NaAlO2, NaAlSiO4) phases, which improves the dissolution in sulfuric acid compared to the oxide phases of metals. The alkali dose was a prominent factor in forming sodium metal oxides. Baking at 600 °C, 1 h, 75 wt% NaOH yielded 60% Al dissolution during water leaching and 82% Ti, 83% Fe, 97% Sc, and 64% Ga dissolution during acid leaching. Scandium dissolution strongly correlates with titanium during leaching, and TEM analysis confirms the apparent enrichment of scandium in the anatase and hematite matrix. Carbonation of water leach solution precipitates alumina and dawsonite (NaAlCO3(OH)2) phases with a product having 85–87% Al2O3 purity and 12 wt% yield. Thermal hydrolysis of acid leach solution generates Ti(OH)4 precipitate which is calcined to obtain anatase with a 95–97% TiO2 and 13 wt% yield. The leach residue contains unreacted hematite and anatase phases having <10% Al2O3 and SiO2 with 6.5 wt% yield. The hydrolyzed solution contains Sc(III) (6 mg/L), Ga(III) (5 mg/L), and Fe(II) (28.4 g/L) ions. The proposed process consumes 0.75 kg NaOH and 1.40 kg H2SO4 per kg red mud and generates 0.14 kg alumina and 0.13 kg titania product with 0.07 kg residue enriched with 48% anatase.