A Novel Process for Al₂O₃ and KCl Production from Mechanically Activated K‑Feldspar via HCl Leaching and Roasting

Abstract

This study investigates the recovery of potassium and aluminum from a Turkish K-feldspar ore via a combined process of intensive milling for mechanical activation, hydrochloric acid (HCl) leaching, and evaporation–roasting for the production of potassium chloride (KCl) and alumina (Al₂O₃). Mechanical activation was performed in a planetary ball mill at a ball-to-ore ratio of 30, 400 rpm for 120 min, which resulted in complete amorphization of microcline and albite, while quartz remained partially crystalline (13.91%) due to its homogeneous Si–O–Si network. The activated ore exhibited a sixfold increase in specific surface area (from 4.68 to 28.05 m²/g) and a significant increase in loss on ignition (LOI) from 0.46% to 2.88%, attributed to the formation of surface hydroxyls (Si–OH, Al–OH), confirmed by FT-IR analysis. Under identical leaching conditions (4 M HCl, 100 °C, 24 h, L/S = 20, 600 rpm), the raw ore yielded only 2.83% Al extraction, whereas the optimally milled sample achieved 87.72%, demonstrating the critical role of mechanical activation. Subsequently, response surface methodology (RSM) was employed to optimize the leaching parameters, yielding optimum conditions of 4.88 M HCl, 100 °C, and 36 h, which resulted in 82.08% Al and 84.20% K recoveries. The negative quadratic coefficient for HCl concentration (C² = −8.65) reflects the decreased solubility of AlCl₃·6H₂O in concentrated HCl due to chloro-aluminum complex formation (AlCl₃(aq), AlCl₄ ^–). The loaded leach solution was directly evaporated and roasted at 1150–1250 °C to separate Al₂O₃ from KCl, which volatilized and condensed as a salt mixture containing 81.3% KCl and 17.3% NaCl; the KCl fraction can be further purified to a high-purity product. Contamination from WC–Co milling media was fully eliminated by changing to sintered alumina media, which also improved Al₂O₃ purity to 94.1% and produced a white product. The silica-rich leach residue (90.87% SiO₂) was valorized into high-purity SiO₂ (97.7%), calcium silicate (CaSiO₃), and silicon carbide (SiC) via biochar-driven carbothermal reduction. Techno-economic analysis of the integrated POTASSIAL process (100 ton/day) indicated marginal baseline economics (IRR ≈ 6.1%, payback ≈ 12 years) but became highly attractive (IRR > 22%, payback <4 years) with a 30% increase in SiC price, while LCA confirmed long-term environmental benefits from closed-loop operation and zero-waste generation. This work demonstrates a sustainable, zero-waste pathway for the simultaneous valorization of K and Al from feldspathic ores, offering a viable alternative for regions lacking soluble potash resources.

1. Introduction

Aluminum is one of the most abundant metals in the Earth’s crust, accounting for approximately 8% of its composition, and is the second most commonly used metal after steel. However, due to its high affinity and reactivity to oxygen, it is rarely found in elemental form in nature. Instead, it is found as part of aluminum compounds in over 270 different minerals. Also, alumina (Al₂O₃), the most common Al-containing compound, is widely used in ceramics, abrasives, refractories, and cement industries due to its desirable physical and chemical properties. It is also used in the production of aluminum metal through the Hall–Héroult process, following the Bayer process. However, bauxite-based alumina production faces significant technological and environmental challenges. These include insufficient bauxite reserves in certain countries, depletion of existing reserves, reactive silica issues, high energy consumption, CO₂ emissions, and challenges in red mud storage and management. Rapid technological advancements have led to an increase in the global consumption of aluminum, with a decrease in high-grade bauxite reserves, thus spurring research efforts toward the utilization of non-bauxite ores as a secondary source such as kaolinite, , alunite, mullite, muscovite, sericite, pyrophyllite, − nepheline syenite, anorthite and coal shales, and fly ash. ,

Extraction of Al₂O₃ from clay minerals is a complex process that involves several steps, including: (i) thermal or mechanical pretreatment for activation, (ii) acid leaching, (iii) solution purification, (iv) aluminum precipitation, and (v) roasting at high temperatures to convert it into alumina (Al₂O₃). The purity of the alumina obtained from this process depends on the purity of the starting clay mineral and the efficiency of the various steps in the process. Various studies exist on obtaining alumina from aluminum-containing minerals using the acid leaching method. ,,− In a study conducted by Erdemoğlu et al., it was concluded that mechanical activation is more efficient than thermal activation in the recovery of alumina from aluminosilicate minerals. Therefore, achieving mechanical activation through intensive milling is considered a viable method for high-efficiency metal recovery from minerals. − In fact, without such mechanical activation, the highly stable crystalline structure of aluminosilicates is practically resistant to acid attack, often resulting in insignificant dissolution (<3% Al extraction); thus, intensive milling acts as a precondition rather than simply an optional pretreatment.

Potassium is one of the essential elements for plant nutrition and development. However, its demand for potassium fertilizers is increasing steadily at 3–3.5% per annum. At present, the predominant source for global potash fertilizer production is soluble K salt ores such as sylvite (KCl), which is the mineral form of potassium chloride (KCl). Unfortunately, soluble K salt resources are limited, constituting less than 1% of global K reserves. According to General Directorate of Mineral Research and Exploration, Türkiye does not have any commercial resources of soluble potassium-bearing rocks such as sylvinite (KCl–NaCl) or carnallite (KCl.MgCl₂·6­(H₂O)). Due to the limited domestic raw material resources in Türkiye, the chemical fertilizer industry depends more than 90% on imports. For example, 85% (6 million tons) of the annual chemical fertilizer consumption of around 7 million tons is imported. Consequently, the recovery of potassium from diverse domestic sources is of paramount importance. Generally, potassium is associated with silicate minerals such as feldspar, mica, biotite, muscovite, and nepheline within the Earth’s crust.

The beneficiation of feldspar ores is commonly carried out through the flotation method. , In the recovery of elements within feldspar minerals, potassium is primarily targeted. Calcium-based compounds are often preferred to enhance potassium recovery. − However, since feldspars can contain a notable amount of aluminum, its recovery has also become an important research area. Nevertheless, studies in this field are more limited compared to potassium recovery. ,

Türkiye has huge resources of silicate minerals in the form of feldspar in which potassium concentration is around 5–10%. Feldspars constitute a significant group of aluminosilicate minerals, typically containing sodium, calcium, and potassium. Their widespread occurrence makes them intriguing subjects for geological and crystallographic studies. Importantly, their ice nucleation properties vary depending on the specific feldspar type. Feldspars are mainly used in ceramic and glass industries as fluxing agents.

However, the chemical properties of K-feldspar are stable due to its fixed Al–Si–O tetrahedral structure, making it difficult for KAlSi₃O₈ to dissociate into free K⁺ ions and dissolve in water or soil. Although numerous methods such as thermal roasting, hydrothermal ion exchange, acid/alkali leaching, and bioleaching have been attempted to extract soluble potassium salts from K-feldspar, these processes require large quantities of expensive raw materials (KCl, Na₂CO₃, NaOH, KOH, HCl, HF, and H₂SO₄) and equipment, leading to high costs for using KAlSi₃O₈.

Studies are being conducted to obtain potassium from silicate minerals as an alternative to KCl production from evaporite potash deposits. Hydrometallurgical leaching of potassium from aluminosilicates is generally unsuccessful because of their complex structure. However, methods have been developed to facilitate KCl recovery through water leaching by applying thermal treatments such as roasting with calcium-containing compounds like CaCl₂. ,, Alyosif et al. investigated the process of feldspar roasting with calcium chloride for potassium chloride production from feldspars, in which mechanical activation was achieved by amorphizing minerals subjected to intensive milling for varying durations, and the potassium recovery obtained from roasting these minerals with CaCl₂ was found to be higher at all temperatures compared to that of the unmilled ore. As a pretreatment method, mechanical activation is defined as a process that enhances the reactivity of a solid without altering its chemical composition. In the process, the mineral is intensively milled to fine sizes, reducing the particle size. Depending on the intensity of the applied mechanical energy during milling, deformation occurs in the crystal structure of the mineral, leading to the formation of an amorphous structure. Simultaneously, new surfaces that have not undergone any chemical reactions are exposed. This structure becomes more active during metallurgical processes such as calcination or leaching, resulting in a decrease in reaction temperature or an increase in process rate.

Response surface methodology (RSM) has proven to be a valuable tool in hydrometallurgical research due to its ability to optimize processes involving multiple operating parameters with a minimal number of experiments. Its advantagessuch as facilitating easy optimization, reducing experimental workload, enabling rapid result acquisition, and allowing statistical evaluationhave led to its widespread use in recent years. In the context of potassium and aluminum recovery from K-feldspar through acid leaching, RSM is expected to be particularly beneficial for systematically investigating the effects and interactions of leaching variables and for efficiently determining the optimum conditions for maximum Al and K solubility.

Previous studies on recovering potassium as KCl from potassium feldspar ores have left the aluminum in the ore as waste because it is bound to various calcium–aluminum silicate phases. This occurs as a result of thermochemical reactions with added calcium during chlorination roasting. However, no other study has been found in which both the potassium and aluminum were taken into an aqueous solution by the acid leaching of K-feldspar, and KCl and Al₂O₃ were recovered as separate products. Furthermore, the conventional downstream purification of acid leachates often relies on selective pH precipitation (e.g., Al­(OH)₃ at pH 4–6), which introduces external reagents such as NaOH, causes co-precipitation of contaminant cations (Fe³⁺, Co²⁺, etc.), and generates large volumes of waste salts.

In contrast, the present work adopts a reagent-free evaporation–roasting route: the loaded chloride leachate is evaporated to recover HCl and co-crystallize AlCl₃·6H₂O and KCl; subsequent roasting transforms the aluminum into solid Al₂O₃ retained in the furnace container, while KCl volatilizes and is separately collected by condensation on a cooled surface.

KCl volatilizes because it exists in the liquid–vapor phase in the temperature range of 906–1418 °C. KCl mass loss begins around 800 °C and continues until approximately 1190 °C, where it is completely consumed. Similar observations have been reported for the roasting of AlCl₃–KCl salt mixtures derived from pyrophyllite ore, where volatilized KCl condensed in the cooler zones of the furnace. , This integrated methodology eliminates the need for costly precipitants, minimizes secondary waste, and enables acid recycling, offering a distinct advantage over conventional purification schemes.

This research aims to investigate the extraction of potassium and aluminum from intensively milled potassium feldspar ore, for which mechanical activation is achieved, through an acid leaching process, followed by roasting performed to obtain Al₂O₃ and KCl. The study focuses especially on optimizing the acid leaching parameters affecting the degree of potassium and aluminum dissolution from mechanically activated potassium feldspar ore into the leaching solution utilized to obtain KCl and Al₂O₃. By systematically comparing the leaching performance of raw versus activated ore and implementing a closed-loop chloride separation strategy, the current investigation provides a comprehensive framework for the simultaneous valorization of both K and Al values from a domestic feldspar resource.

2. Materials and Methods

2.1. Materials

The K-feldspar sample used in this study was obtained from Kale Seramik Company, Türkiye, as a flotation concentrate from the beneficiation of feldspar ore. The raw material was subjected to intensive milling in a planetary ball mill based on a Taguchi experimental design. The milled products were characterized using X-ray diffraction (XRD), particle size distribution (PSD), and BET surface area analysis; the results of this milling study have been reported in detail elsewhere. Based on the characterization and leaching tests conducted in the present work, the milled K-feldspar that exhibited the highest Al and K recoveries was selected for subsequent leaching experiments. A detailed characterization of both raw and activated samples is provided in the Results and Discussion section.

2.2. Intensive Milling

The K-feldspar ore was intensively milled to achieve mechanical activation by using a Fritsch Pulverisette 6 Classic Line model planetary mono mill (Idar-Oberstein, Germany). The feldspar ore was fed into a 250 cm³ milling bowl made of tungsten carbide (WC) with a chemical composition of 88% WC and 12% cobalt (Co). The bowl contained 50 pieces of 10 mm diameter balls made of the same material with the bowl. The ball-to-ore ratio was kept constant at 30 by weight. The ore was dry-milled for 120 min at a 400 rpm rotational mill speed. To reduce overheating of the mill, each 10 min of milling was followed by 20 min of pause. The milling tests were carried out in an air atmosphere and at room temperature.

2.3. Acid Leaching

For each experiment, 12.5 g of sample was leached in 250 mL of HCl solution using a 500 mL two-necked borosilicate glass reactor equipped with a water-cooled condenser. The reactor containing mechanically activated K-feldspar and HCl solution was placed in a stirring heater mantle, and the temperature of the reactor was controlled by a programmable controller of the mantle. A magnetic stirrer was used to mix the slurry while keeping the stirring speed constant at 600 rpm. At the end of leaching, the slurry was vacuum-filtered to separate the loaded leach solution from undissolved material, which was then washed three times with diluted HCl acid solution. The filtrate and washings were combined to have a constant volume of 500 mL in a volumetric flask to obtain a stock-loaded leach solution, subjected to inductively coupled plasma-optical emission spectroscopy (ICP-OES) analysis.

2.4. Evaporation and Roasting for Al₂O₃ and KCl Recovery

The clarified loaded leach solution obtained under the optimized leaching conditions was subjected to direct evaporation without any prior neutralization or pH adjustment. The solution was transferred to a glass reactor and heated to its boiling point (∼100 °C) under atmospheric pressure. The evolved vapors, consisting of water steam and gaseous HCl, were passed through a water-cooled condenser, and the recovered acidic condensate was collected for potential reuse as a leachant. As evaporation proceeded, the dissolved salts reached supersaturation, leading to the co-crystallization of hydrated AlCl₃ and KCl at the bottom of the evaporation reactor.

The resulting mixed salt lump was then transferred to a high-purity alumina crucible and placed inside a horizontal three-zone tube furnace (Protherm PZF12/105/750, capable of reaching 1200 °C) for roasting. To collect the volatilized salts, a borosilicate glass tube equipped with inlet and outlet caps was placed inside the mullite tube of the furnace. The salt mixture was roasted at 1150 °C. During roasting, the aluminum species remained in the crucible as insoluble alumina. In contrast, the volatile KCl rerouted and was transported by the thermal gradient to the cooler region of the borosilicate tube, where it condensed and was collected as a solid crystalline deposit. After the volatile KCl fraction was removed, the remaining alumina in the crucible was further calcined at 1200 and 1250 °C using a muffle furnace (Protherm PLF120/5, capable of reaching 1300 °C) to obtain the final high-purity Al₂O₃ product. This evaporation–roasting scheme has been successfully applied in our previous works ,, and is the subject of a patent application.

This temperature-driven separation enabled the simultaneous recovery of both Al₂O₃ and KCl without the addition of any external chemical reagents, while the condensed acidic distillate allowed for the recycling of a portion of the leachant.

2.5. Characterization Techniques

The crystal structure of the samples was examined via X-ray diffraction (XRD) using an automated Rigaku RadB-type diffractometer (Tokyo, Japan). The radiation applied was CuKα (λ = 1.5405 Å) from a long fine-focus Cu tube, operating at 40 kV and 40 mA. The samples were measured in the 2θ range of 2–80° with 0.02° steps at a rate of 2° per min. The Jade 5.0 program was used for phase identification by matching the ICDD PDF database with the measured samples. Degree of amorphization was calculated using the crystallinity index of the minerals, which is a measure of the relative amount of crystalline material in a sample. It is calculated by dividing the area of the crystalline peaks in an XRD pattern by the total area of the pattern. The higher the crystallinity index (CI), the more crystalline the sample. The degree of crystallinity and amorphization is formulated in eqs and , respectively

$$\mathrm{T}\mathrm{h}\mathrm{e}\mathrm{c}\mathrm{r}\mathrm{y}\mathrm{s}\mathrm{t}\mathrm{a}\mathrm{l}\mathrm{l}\mathrm{i}\mathrm{n}\mathrm{i}\mathrm{t}\mathrm{y}\mathrm{i}\mathrm{n}\mathrm{d}\mathrm{e}\mathrm{x}(\mathrm{C}\mathrm{I}),\%=(U_0{I}_x)/(I_0{U}_x)\times 100$$ 1

$$\mathrm{T}\mathrm{h}\mathrm{e}\mathrm{d}\mathrm{e}\mathrm{g}\mathrm{r}\mathrm{e}\mathrm{e}\mathrm{o}\mathrm{f}\mathrm{a}\mathrm{m}\mathrm{o}\mathrm{r}\mathrm{p}\mathrm{h}\mathrm{i}\mathrm{z}\mathrm{a}\mathrm{t}\mathrm{i}\mathrm{o}\mathrm{n}(\mathrm{D}\mathrm{A}),\%=(100-\mathrm{C}\mathrm{I})$$ 2

where U ₀ and U _x denote the backgrounds of unmilled sample and milled sample, while I ₀ and I _x are the integral intensities of diffraction lines of the mineral in the unmilled sample and milled sample, respectively. Accordingly, if I _x = I ₀, it is interpreted that the mineral is 100% XRD-crystalline or 0% XRD-amorphous with respect to the crystallographic plane (hkl) to which the peak belongs.

The chemical composition of the ore was determined by Inductively Coupled Plasma-Atomic Emission Spectroscopy (ICP-AES) in ALS Global Geochemistry Lab (Türkiye).

Al and K concentrations in the leach solution were determined by Flame Atomic Absorption Spectrometry (FAAS) at İnönü University, İBTAM Laboratories. The elemental content of the loaded leach solution obtained under optimum conditions was determined by the Inductively Coupled Plasma-Optic Emission Spectroscopy (ICP-OES) method (Middle East Technical University Central Laboratory, Türkiye).

The particle size distribution (PSD) of the samples was determined using a Malvern Mastersizer 2000 particle size analyzer (Malvern Instruments Ltd., UK). Nearly 0.5 g of sample was stirred with distilled water at 2000 rpm in a 1 L cell integrated into the device, and measurements were performed on particles in an aqueous medium. All of the particle size distribution data were automatically obtained by using Mastersizer 2000 software.

The specific surface area of the powders was determined by the BET method using the low-temperature N₂ adsorption technique using a NOVA 1200e pore size analyzer (Quantachrome Instruments, UK). Thermal analysis of all samples was performed using a Shimadzu TGA-50 thermogravimetric analyzer. The chemical composition of the leach residue was determined by the XRF method using a PanAnalytical Epsilon 4 instrument.

Fourier transform infrared spectroscopy (FT-IR) analysis of the raw and activated feldspar ore was performed using a PerkinElmer Spectrum One model spectrometer.

3. Results and Discussion

3.1. Characterization of Raw and Activated Feldspar Ore

Figure presents the XRD patterns of raw and milled K-feldspar ore obtained through intensive milling in a planetary ball mill. When the XRD pattern of the milled sample was examined, it was observed that the peak intensities of all minerals typically declined and the peaks belonging to feldspar minerals almost entirely disappeared. The remaining XRD peaks observed in the patterns after mechanical activation were attributed to quartz. The most intense peak associated with quartz, appearing at 26.56° 2θ (CuKα), shows a slight decrease in intensity. This suggests that the crystal structure of quartz within the ore is also distorted to some extent due to intensive milling. Petrov and Milošević explain that varying resistances to mechanical activation are due to the different crystal lattice types. In this context, quartz’s higher resistance to amorphization (only 13.91%, as shown in Table ) is not solely due to its higher Mohs hardness (7 versus 6 for feldspar) but due to the homogeneous, three-dimensional Si–O–Si network of quartz, where bond breakage requires cooperative failure along specific crystallographic directions. In contrast, feldspars contain weaker Al–O bonds and charge-balancing cations (e.g. K⁺, Na⁺), which introduce local lattice distortions and internal strain, making the framework less durable against high-energy milling. This mechanistic dissimilarity clarifies why microcline and albite reach 100% amorphization while quartz retains partial crystallinity. , The structural changes implied by the XRD patterns are further corroborated by FT-IR spectroscopic analysis, which provides direct molecular-level evidence for bond weakening and lattice distortion (discussed below in this section).

1.

XRD patterns of the raw and mechanically activated feldspar (Mc: Microcline, Ab: Albite, Qz: Quartz, WC: Tungsten carbide).

2. Particle Size and Amorphization Parameters of Raw and Activated Feldspar Samples.

feldspar ore	particle size parameter, μm			S BET, m2/g	degree of amorphization, %
	d10	d50	d90		microcline	albite	quartz
raw	121.42	293.11	518.30	4.68	0	0	0
activated	0.95	20.40	218.20	28.05	100	100	13.91

The chemical composition of the raw and activated K-feldspar ore sample is given in Table . A relative decrease was observed in the activated product’s silica (65.03% SiO₂) and alumina (15.08% Al₂O₃) contents when compared to the raw ore. However, a more significant result was the presence of tungsten (1.59% WO₃) and cobalt (0.12% Co₃O₄) in the activated material. The presence of these elements is a clear indication of contamination originating from the milling media. Another result was the significant difference in loss on ignition (LOI) values. The raw ore had an LOI of 0.46%, while the activated ore’s LOI was determined to be 2.88%. The origin of this LOI increase is further investigated by thermal analysis and FT-IR spectroscopy in the following sections.

1. Chemical Compositions of the Raw and Activated Feldspar Samples.

main oxide, %	SiO2	Al2O3	Fe2O3	CaO	MgO	Na2O	K2O	TiO2	BaO	WO3	Co3O4	LOI
raw	69.23	16.69	0.14	0.34	0.03	1.80	10.94	0.29	0.11	-	-	0.46
activated	65.03	15.08	0.19	0.35	0.01	1.86	10.05	0.06	0.09	1.59	0.12	2.88

^aLoss on ignition at 1000 °C.

Table shows data on the amorphization degrees and particle size parameters for both raw and activated samples. This table clearly demonstrates the effects of the intensive milling process on the structural properties of the materials. Particle size measurements showed that d₉₀, d₅₀, and d₁₀ sizes of raw ore are 518, 293, and 121 μm, respectively. But, for activated feldspar, these values shifted to 218, 20, and 0.9 μm, respectively. Additionally, mechanical activation led to six times increase in the specific surface area from 4.68 to 28.05 m²/g. The degree of amorphization calculated according to eq clearly reveals the change in mineral crystal structural characteristics. Table shows the degree of amorphization of microcline, albite, and quartz, being the major minerals in the ore. These minerals in the ore are subjected to amorphization by milling just for 120 min. For quartz, the amorphization degree is about 13.91% at the end of 120 min of milling, while it is 100% for microcline and albite, respectively. Unlike quartz mineral, the amorphization degrees of microcline and albite decrease with further milling.

Scanning Electron Microscopy–Energy Dispersive X-ray (SEM–EDX) analysis was conducted to investigate the surface morphology of the raw and activated feldspar samples. The surface characteristics of the ore samples are shown in Figure . As can be seen from the SEM image of raw ore (i.e., unmilled) with −212 μm particle size (Figure a), there are many particles with different sizes and shapes including coarse ones with sharp edges and fine ones with smooth surfaces. The wide variation in shapes is because of the presence of hard minerals such as microcline and quartz. Furthermore, the coarser particles were characterized by a more homogeneous internal structure and absence of cracks. In contrast, the mechanically activated feldspar displays smaller particles, irregular surface morphology, and prominent fractures with sharp edges. Moreover, mechanical activation led to an increase in surface area, resulting in the agglomeration of particles (Figure b). When evaluated in terms of chemical composition, it was determined that the SiO₂ content decreased in the raw sample, while the proportions of K and Al increased. This change can be considered an indication of structural transformations caused by mechanical activation within the feldspar structure. The significant increase in potassium content particularly suggests that feldspar’s structural integrity was compromised, leading to the release of potassium ions from its surface. On the other hand, EDX analysis of the activated sample, which was obtained through milling in a tungsten carbide medium, revealed a dominance of Si and Al elements, alongside the distinct appearance of new element peaks such as W and Co. This indicates contamination of the sample by the milling media due to abrasion, especially during prolonged milling times.

2.

SEM–EDX microphotographs of raw (a) and activated (b) feldspar samples.

The apparent discrepancy between the EDX (increased K/Al ratios relative to Si) and the ICP-AES data, regarding relative decrease in Al₂O₃ and K₂O (Table ), is explained by the dilution effect of the added contaminants and LOI. By recalculating the activated sample’s composition after excluding LOI, WO₃, and Co₃O₄ (i.e., normalizing to the original mineral oxide basis only), the values become: SiO₂ = 70.14%, Al₂O₃ = 16.27%, and K₂O = 10.84%. Compared to the raw feldspar on the same dry, contaminant-free basis (SiO₂ = 69.55%, Al₂O₃ = 16.77%, K₂O = 10.99%), the differences are only −0.50% for Al₂O₃ and −0.15% for K₂O. Thus, the apparent decrease in bulk Al and K is largely a dilution effect, while the EDX surface observations reflect amorphization-induced surface enrichment. The management of WC and Co contamination is discussed in detail in Section , along with the corresponding purification results.

Thermal analysis (TGA/DTA) was performed on the raw and activated ore sample, and the results are presented in Figure . According to TGA curve for the raw ore, mass loss began at 384 °C, reaching approximately 1% at 760 °C, with a maximum loss of 1.06% at 1000 °C. It clearly shows that raw ore exhibits very high thermal stability over the entire temperature range. A slight mass loss of 1–2% is observed at high temperatures (between 850 and 1000 °C). This suggests that the sample contains little or no volatile components (such as moisture, organic matter, and carbonate) and can be a high-purity feldspar ore. Feldspar minerals generally do not show significant mass loss in this temperature range. This mass loss is attributed to the dehydroxylation of muscovite, which is present in the the ore as a minor gangue mineral. For pure muscovite, the mass loss upon heating to 1000 °C is approximately 3.82%. As for the activated sample, the TGA curve shows a much more significant mass loss compared to the raw ore. It is seen that at 1000 °C, the sample has lost about 5–6% of its initial mass. The high percentage of mass loss clearly indicates that this sample contains significant amounts of thermally decomposable components such as moisture (i.e., OH groups).

3.

TGA and DTA curves of the raw and activated K-feldspar.

The exothermic peak at 710.45 °C was assessed, as direct decomposition or phase transitions are not typically expected for feldspars in this temperature range. However, if impurities like quartz are present along with feldspar in the sample, the α-quartz to β-quartz transformation is characterized by a weak endothermic peak. Smaller exothermic peaks at 904.12 °C and 960.64 °C were assessed as these peaks similarly could indicate high-temperature phase transitions, the onset of sintering, or the beginning of melting of minor components or impurities in the ore. The peak at 960.64 °C, in particular, is close to the typical melting temperature for feldspars, though this is usually an endothermic event. Significantly, the sharp and broad exothermic peak observed above 850 °C (reaching a maximum at 950 °C) is absent in the raw ore but pronounced in the activated sample. This exotherm is characteristic of the crystallization of a metastable Al-rich spinel phase (γ-Al₂O₃-type) or mullite from the highly amorphous aluminosilicate matrix generated by mechanical activation. , This crystallization event is sequential to the volatile loss: the 2.88% LOI reflects the release of adsorbed and structurally bound water/hydroxyl groups from the defective, highly porous activated structure up to about 850 °C, after which the dehydrated amorphous phase reorganizes into a crystalline spinel/mullite phase, producing the observed exothermic DTA peak without further significant mass change.

The FT-IR spectra (Figure ) provide spectroscopic confirmation of the structural amorphization observed by XRD (Figure ) and confirmation for the origin of the increased LOI in the activated sample. FT-IR analysis was performed on both samples. The raw K-feldspar exhibited no absorption bands in the characteristic O–H stretching region (3200–3700 cm^–1). In contrast, the activated sample revealed a distinct new broad band centered at approximately 3400 cm^–1, along with a minor bending band near 1640 cm^–1, which are unequivocally assignable to O–H stretching and H–O–H bending vibrations, respectively. This confirms the presence of both chemisorbed surface hydroxyls (Si–OH, Al–OH) and physisorbed molecular water on the freshly fractured surfaces. The broadness of the 3400 cm^–1 band indicates extensive hydrogen bonding, typical of water clusters on silicate surfaces. These spectral changes, together with the increased mass loss observed in TGA, provide direct evidence that mechanical activation induces significant surface hydroxylation and water adsorption, which account for the LOI increase from 0.46% to 2.88%. Results indicate significant changes on the surface properties of the particle during the milling process. As is known, milling of minerals reduces particle size while substantially increasing the specific surface area. These newly formed surfaces have unsaturated bonds or unsaturated atoms in terms of coordination, which are the result of the breaking of bonds in the crystal lattice. These reactive sites easily interact with water molecules in the atmosphere to form hydroxyl groups. This process is called hydration or hydroxylation. Furthermore, these spectral changes are fully consistent with the XRD-derived amorphization degrees (Table ): the complete loss of crystallinity in microcline and albite corresponds to the disappearance of sharp bands and the broadening/shifting of the Si–O–Si stretching modes, while the partial crystallinity of quartz is reflected in the less pronounced spectral changes. Thus, the FT-IR data provide direct molecular-level evidence that mechanical activation disrupts the Al–O–Si linkages preferentially, weakening the framework and rendering it vulnerable to acid attack.

4.

FT-IR spectra of raw and mechanically activated K-feldspar ores in the 4000–400 cm^–1 range (Abbreviations: ν, stretching; δ, bending; as, asymmetric; s, symmetric).

3.2. Leaching of Silicate Minerals for K and Al Extraction

Aluminum silicate minerals are known for their low solubility and slow dissolution. The crystal lattice of these minerals is characterized by strong covalent or ionic Si–O bonds forming a tetrahedral layer and ionic Al–O bonds. Therefore, during the dissolution of clay minerals, a dominant transport of aluminum and deposition of silicon are observed. In boiling concentrated acids, the H⁺ ions from the acid react with the Al–O bonds into replacing the Al³⁺ ions. Meanwhile, nonstable orthosilicic acid groups Si­(OH)₄ or H₄SiO₄ in hot acid solutions are formed. Eqs – represent the dissolution reactions of various aluminum silicate minerals in acidic conditions:

$${\mathrm{K}\mathrm{A}\mathrm{l}\mathrm{S}\mathrm{i}}_3{\mathrm{O}}_8(\mathrm{M}\mathrm{i}\mathrm{c}\mathrm{r}\mathrm{o}\mathrm{c}\mathrm{l}\mathrm{i}\mathrm{n}\mathrm{e})+4{\mathrm{H}}^{+}\to {\mathrm{K}}^{+}+{\mathrm{A}\mathrm{l}}^{3+}+3{\mathrm{S}\mathrm{i}\mathrm{O}}_2(\mathrm{a}\mathrm{q})+2{\mathrm{H}}_2\mathrm{O}$$ 3

$${\mathrm{N}\mathrm{a}\mathrm{A}\mathrm{l}\mathrm{S}\mathrm{i}}_3{\mathrm{O}}_8(\mathrm{A}\mathrm{l}\mathrm{b}\mathrm{i}\mathrm{t}\mathrm{e})+4{\mathrm{H}}^{+}\to {\mathrm{N}\mathrm{a}}^{+}+{\mathrm{A}\mathrm{l}}^{3+}+3{\mathrm{S}\mathrm{i}\mathrm{O}}_2(\mathrm{a}\mathrm{q})+2{\mathrm{H}}_2\mathrm{O}$$ 4

$${\mathrm{A}\mathrm{l}}_2{\mathrm{S}\mathrm{i}}_2{\mathrm{O}}_5{(\mathrm{O}\mathrm{H})}_4(\mathrm{K}\mathrm{a}\mathrm{o}\mathrm{l}\mathrm{i}\mathrm{n}\mathrm{i}\mathrm{t}\mathrm{e})+6{\mathrm{H}}^{+}\to 2{\mathrm{A}\mathrm{l}}^{3+}+2{\mathrm{S}\mathrm{i}\mathrm{O}}_2(\mathrm{a}\mathrm{q})+5{\mathrm{H}}_2\mathrm{O}$$ 5

$${\mathrm{K}\mathrm{A}\mathrm{l}}_3{\mathrm{S}\mathrm{i}}_3{\mathrm{O}}_{10}{(\mathrm{O}\mathrm{H})}_2(\mathrm{M}\mathrm{u}\mathrm{s}\mathrm{c}\mathrm{o}\mathrm{v}\mathrm{i}\mathrm{t}\mathrm{e})+10{\mathrm{H}}^{+}\to 3{\mathrm{A}\mathrm{l}}^{3+}+{\mathrm{K}}^{+}+3{\mathrm{S}\mathrm{i}\mathrm{O}}_2(\mathrm{a}\mathrm{q})+6{\mathrm{H}}_2\mathrm{O}$$ 6

$${\mathrm{A}\mathrm{l}}_2{\mathrm{S}\mathrm{i}}_4{\mathrm{O}}_{10}{(\mathrm{O}\mathrm{H})}_2(\mathrm{P}\mathrm{y}\mathrm{r}\mathrm{o}\mathrm{p}\mathrm{h}\mathrm{y}\mathrm{l}\mathrm{l}\mathrm{i}\mathrm{t}\mathrm{e})+6{\mathrm{H}}^{+}\to 2{\mathrm{A}\mathrm{l}}^{3+}+4{\mathrm{S}\mathrm{i}\mathrm{O}}_2(\mathrm{a}\mathrm{q})+4{\mathrm{H}}_2\mathrm{O}$$ 7

These reactions represent the dissolution of silicate minerals containing aluminum and alkali metals under an acidic environment containing H⁺ ions. When an acid such as HCl is present, feldspars (e.g., microcline and albite) and clay minerals (e.g., kaolinite, muscovite, and pyrophyllite) react with protons and eventually dissolve. The reactions release structural alkali ions (K⁺ and Na⁺) and Al³⁺ ions, while the silicon content is generally released as dissolved SiO₂ (silica gel). These reactions form the basis for recovering valuable compounds, such as Al₂O₃ and KCl, using an acidic leaching method.

As for the selection of acid type for leaching, Bazin et al. produced Al₂O₃ from aluminum chloride salt by applying H₂SO₄ leaching, followed by HCl leaching. Due to the relatively high decomposition temperature of Al₂(SO₄)₃, the sulfated solution is chlorinated with HCl to obtain Al₂O₃ at lower decomposition temperatures. Erdemoğlu et al. produced Al₂O₃ with high recovery by direct HCl leaching of mechanically activated pyrophyllite ore. Moreover, TiO₂ and SiO₂ are practically insoluble in HCl. Therefore, it reduces the need for the separation of Ti from the leachate. It is worth noting that in the present integrated flowsheet, HCl was chosen not merely as a lixiviant but as a strategic reagent: the generation of chloride salts (KCl and AlCl₃) is essential for the subsequent evaporation–roasting step, where KCl volatilizes at high temperatures while Al remains as Al₂O₃ in the solid residuea separation that is not possible with sulfate systems. This process integration rationale guided our acid selection. XRD analysis (Figure ) indicates that the feldspathic ore is predominantly composed of microcline (KAlSi₃O₈), along with minor concentrations of albite (NaAlSi₃O₈) and muscovite (KAl₂(AlSi₃O₁₀)­(OH)₂). Following mechanical activation, the ore is leached in an acidic solution, causing Al³⁺ and K⁺ ions to be transferred into the leachate. When HCl was selected as the leaching agent, the dissolved ions precipitate as their corresponding chlorine salts (i.e., KCl and AlCl₃) in the medium.

To quantitatively evaluate the effectiveness of the mechanical activation step, leaching experiments were first performed on both the raw feldspar and the milled samples obtained under different milling conditions while keeping the leaching conditions fixed: 4 M HCl, 24 h, ∼100 °C (boiling point), liquid-to-solid ratio (S/L) of 20, and stirring speed of 600 rpm. Under these fixed conditions, the raw feldspar yielded only 2.83% aluminum extraction. In contrast, the highest aluminum extraction (82.08%) was achieved after intensive milling of the ore at a ball-to-ore ratio of 30, a mill speed of 400 rpm, and a milling time of 120 min. This nearly 30-fold increase unequivocally demonstrates that intensive milling is not merely a pretreatment but a prerequisite for rendering the feldspar structure sufficiently ready to acid leaching. Based on these screening tests, the optimally milled feldspar was selected for all subsequent RSM optimization studies.

3.3. Optimization of the Leaching Parameters

The Box–Behnken experimental design model, one of the RSM methods, was used in the experimental studies. HCl concentration, leaching temperature, and leaching time were defined as independent variables, while aluminum and potassium dissolution was defined as response values. Other parameters affecting the leaching process, such as the stirring speed (600 rpm) and solution-to-solid ratio, were kept constant. The solid-to-liquid ratio was fixed at 1:20 (20 mL/g) throughout all of the experiments. This value was selected after preliminary scoping tests to ensure a sufficient acid excess relative to the solid feed, thereby preventing acid depletion during the prolonged leaching period. Additionally, this pulp density was low enough to avoid the formation of viscous silica gels, which are known to impede mass transfer and complicate filtration when leaching aluminosilicates at high solid densities. To validate this choice, a separate test was performed at a higher S/L ratio of 1:40 (10 g in 400 mL) under otherwise identical conditions; the aluminum recovery (88.2%) was only marginally higher than that achieved at 1:20 (82.08%), while doubling the liquid volume would lead to increased water/acid consumption, larger reactor volumes, and higher heating costs. Thus, the 1:20 ratio was retained as the optimal compromise between leaching efficiency and process economy. Table presents the experimental ranges and levels of the independent variables.

3. Experimental Levels of Independent Variables.

code and the name of the factor	unit	low	high
A: leaching temperature	°C	30	100
B: leaching time	hour	6	36
C: HCl concentration	M	1	7

Based on the 17 tests designed using the variable levels in Table , Al and K recoveries were calculated by utilizing the loaded leach solution data, and the results are presented in Table .

4. Experimental Design and Results.

order	A	B	C	Al recovery, %	K recovery, %
1	100	6	4	60.94	65.35
2	65	6	4	36.87	44.69
3	65	21	4	56.99	59.26
4	100	21	7	66.78	76.10
5	100	21	1	54.70	59.12
6	65	21	4	51.35	57.47
7	30	21	7	21.20	27.36
8	30	21	1	18.21	27.67
9	65	36	7	60.14	62.31
10	30	36	4	32.92	35.57
11	100	36	4	77.36	69.35
12	65	21	4	62.98	59.35
13	30	6	4	21.46	24.95
14	65	21	4	61.46	58.76
15	65	36	1	57.57	47.89
16	65	6	1	41.87	45.22
17	65	21	4	59.57	58.95

Additionally, using the obtained data, the complex interactions between each parameter were revealed using an experimental design program and presented using an Analysis of Variance (ANOVA) table. This provided information about the interactions. The statistical significance of the effect of independent variables on the results was determined by evaluating the p-values in the ANOVA tables prepared separately for Al and K. A p-value of ≤0.05 indicated that the independent variables had a significant effect on the results. ANOVA values were used to strengthen the statistical validation through the correlation coefficient (R²), one of the parameters indicating the fit of the second-order model.

In the experimental design, used to examine the effect of variables on aluminum dissolution, the Adj-R² value was observed to be (93.66), and the R² value was (97.23). Similarly, in the study conducted for potassium, the Adj-R² value was observed to be (99.28), and the R² value was (99.69). The parameters affecting the results are shown in Table for aluminum and Table for potassium, according to ANOVA values. According to the ANOVA results, the effective parameters for Al dissolution are A, B, A², and C². Furthermore, it can be said that changing the leaching temperature (A²) and HCl concentration (C²) has an exponential effect on the results. According to the ANOVA results, it can be said that all parameters are effective in K dissolution, and all parameters also have an exponential effect on the results. The experimental results are as shown in the model equations for aluminum (eq ) and potassium (eq ) extraction percentages.

$$\mathrm{A}\mathrm{l}\mathrm{R}\mathrm{e}\mathrm{c}\mathrm{o}\mathrm{v}\mathrm{e}\mathrm{r}\mathrm{y}=+58.47+{20.75}^{*}\mathrm{A}+{8.36}^{*}\mathrm{B}+{1.58}^{*}\mathrm{C}+{1.24}^{*}\mathrm{A}\mathrm{B}+{2.27}^{*}\mathrm{A}\mathrm{C}+{1.89}^{*}\mathrm{B}\mathrm{C}-{9.60}^{*}{\mathrm{A}}^2-{0.71}^{*}{\mathrm{B}}^2-{8.65}^{*}{\mathrm{C}}^2$$ 8

$$\mathrm{K}\mathrm{R}\mathrm{e}\mathrm{c}\mathrm{o}\mathrm{v}\mathrm{e}\mathrm{r}\mathrm{y}=+58.76+{19.30}^{*}\mathrm{A}+{4.36}^{*}\mathrm{B}+{3.82}^{*}\mathrm{C}-{1.66}^{*}\mathrm{A}\mathrm{B}+{4.32}^{*}\mathrm{A}\mathrm{C}+{3.74}^{*}\mathrm{B}\mathrm{C}-{6.21}^{*}{\mathrm{A}}^2-{3.74}^{*}{\mathrm{B}}^2-{4.99}^{*}{\mathrm{C}}^2$$ 9

5. ANOVA Values for Al Recovery .

source	sum of squares	degree of freedom	mean square	F value	p value
model	4818.03	9	535.34	27.26	0.0001
A	3444.09	1	3444.09	175.39	<0.0001
B	558.62	1	558.62	28.45	0.0011
C	19.97	1	19.97	1.02	0.3468
AB	6.15	1	6.15	0.31	0.5932
AC	20.66	1	20.66	1.05	0.3392
BC	14.33	1	14.33	0.73	0.4213
A	387.64	1	387.64	19.74	0.0030
B	2.09	1	2.09	0.11	0.7536
C	315.22	1	315.22	16.05	0.0051
residual	137.46	7	19.64
lack of fit	54.08	3	18.03	0.86	0.5287
pure error	83.37	4	20.84
cor total	4955.49	16

^aR = 0.9723; Adj R² = 0.9366.

6. ANOVA Values for K Recovery .

source	sum of squares	degree of freedom	mean square	F value	p value
model	3751.66	9	416.85	247.33	<0.0001
A	2978.76	1	2978.76	1767.37	<0.0001
B	152.34	1	152.34	90.39	<0.0001
C	116.74	1	116.74	69.26	<0.0001
AB	10.96	1	10.96	6.50	0.0381
AC	74.74	1	74.74	44.34	0.0003
BC	55.88	1	55.88	33.15	0.0007
A	162.32	1	162.32	96.31	<0.0001
B	59.02	1	59.02	35.02	0.0006
C	104.70	1	104.70	62.12	0.0001
residual	11.80	7	1.69
lack of fit	9.50	3	3.17	5.51	0.0663
pure error	2.30	4	0.57
cor total	3763.46	16

^aR = 0.9969; Adj R² = 0.9928.

The Normal Plot of Residuals graph, used to test whether the residuals in the regression model conform to a normal distribution and to determine the extent to which the model meets the assumption, and the Predicted vs. Actual data graphs, used to determine how well the model predicted, are presented in Figure for aluminum and Figure for potassium. An examination of the normal plot of residual graphs reveals that the data points for both aluminum and potassium are generally aligned evenly around the red line and that the residuals are normally distributed, meaning that the regression model largely meets this assumption. An examination of the Predicted vs Actual data graphs reveals that the points are largely clustered around the diagonal line, meaning that the model predictions are generally accurate and that there is a generally positive and strong relationship between predicted and actual values. ,

5.

Normal Plot of Residuals and Predicted vs. Actual Data for aluminum.

6.

Normal Plot of Residuals and Predicted vs. Actual data for potassium.

Figure shows the response surface plots of HCl concentration, leaching temperature, and leaching time for aluminum dissolution efficiency. Aluminum dissolution is highest at high leaching temperatures (near boiling point), increasing with the leaching time. Similarly, increasing the HCl concentration leads to higher aluminum recovery. However, due to the low solubility of AlCl₃ in concentrated HCl, aluminum recovery decreases as the HCl concentration increases. Similar results were reported in the studies by Aydemir and Erdemoğlu and Uysal. The negative quadratic coefficient for C² in eq quantitatively captures this optimum. The decline beyond approximately 5 M HCl is attributed to well-documented speciation effects in concentrated chloride media. As the HCl concentration increases, the activity of Cl^– ions rises, shifting the equilibrium from hydrated Al³⁺(aq) toward neutral AlCl₃(aq) and anionic AlCl₄ ^– complexes. , The observed decline beyond ∼5 M HCl can be qualitatively described by the reduced solubility of AlCl₃·6H₂O in concentrated HCl, as estimated from the thermodynamic data for the H⁺–Al³⁺–Cl^––H₂O system. These uncharged or negatively charged species exhibit significantly lower solubility in the low-water-activity, high-ionic-strength medium, leading to their co-crystallization or “salting-out” from the solution phase. Although this phenomenon reduces the measured dissolved Al concentration, it does not necessarily indicate incomplete leaching; rather, it reflects the thermodynamic limitations of the AlCl₃–HCl–H₂O system at extreme acidities.

7.

Effect of parameter interactions on Al dissolution.

Figure shows the response surface plots of HCl concentration, leaching temperature, and leaching time for potassium recoveries. Similarly, potassium dissolution is high at high leaching temperatures (nearly boiling point) and increases with time. Furthermore, an increase in the HCl concentration leads to increased potassium dissolution efficiency. On the basis of the experimental results, the optimization criterion was established to identify the conditions under which the solubility of Al and K is maximized.

8.

Effect of parameter interactions on K dissolution.

Additionally, in order to more clearly monitor the effects of all parameters on aluminum and potassium dissolution efficiencies, one-factor graphs are presented in Figure for aluminum and in Figure for potassium.

9.

Graphs for one factor affecting the aluminum dissolution.

10.

Graphs for one factor affecting the potassium dissolution.

Using the regression equation, three of the optimal leaching conditions determined on the basis of the optimal process conditions that would maximize the response variables (Al and K) were also tested, and the results were found to be very close to the predicted values. The results demonstrated that the developed model has a reliable predictive capacity for the process. The results of the experimental studies conducted under the optimal conditions are presented in Table .

7. Parameter Values Studied at Optimum Solution Points.

parameter	solution point 1	solution point 2	solution point 3
leaching temperature, °C	100	100	100
HCl concentration, M	4882	5537	5214
leaching time, h	35,934	35,103	34,242
Al recovery, %	82,08	79,53	77,70
K recovery, %	84,20	80,38	77,02

The optimum leaching conditions predicted by the model were determined as follows: HCl concentration of 4.88 M, leaching time of 36 h, and leaching temperature of 100 °C (near boiling temperature). Under these optimized conditions, the dissolution efficiencies of Al and K were determined to be 82.08% and 84.20%, respectively. It is acknowledged that the optimal leaching time of 36 h is relatively long compared to typical industrial practice. However, the ANOVA results (Tables and ) show that temperature (A) and its quadratic term (A²) exert the strongest influence on both Al and K recoveries (p < 0.0001), while the time factor (B) has a comparatively smaller coefficient. The results suggest that the leaching process follows a mixed kinetic regime: rapid chemical dissolution dominates the early stages, followed by a slower diffusion-controlled stage limited by proton transport through the evolving silica-rich leached layer. This transition from chemical to diffusion control is consistent with recent studies on aluminosilicate dissolution kinetics, which demonstrate that the formation of a silica-enriched surface layer progressively hinders acid access to the unreacted mineral core. − This kinetic behavior implies that process intensification strategiessuch as higher temperature under pressure or a two-stage acid-addition schemecould potentially achieve equivalent recoveries within significantly shorter times (e.g., 4–8 h). Such alternative intensification pathways will be explored in future work to enhance the industrial scalability of the proposed leaching process.

It is noteworthy that the Al recovery obtained under the RSM-optimized leaching conditions (82.08%) is to some extent lower than the 87.72% Al recovery achieved during the preliminary screening of milling conditions (fixed leaching: 4 M HCl, 24 h). This apparent reduction is attributed to two main factors. First, the screening stage focused exclusively on maximizing Al extraction as a proxy for assessing the effectiveness of mechanical activation, whereas the RSM optimization was designed to simultaneously maximize both Al and K recoveries. The optimized conditions (4.88 M HCl, 36 h) represent a statistical compromise that significantly enhances K recovery to 84.20% while incurring only a minor trade-off in Al recovery. Second, the extended leaching time (36 h vs. 24 h) promotes the development of a silica-enriched leached layer on the particle surfaces, which hinders the outward diffusion of Al³⁺ ions (consistent with the mixed kinetic regime), while allowing K⁺a smaller and more mobile cationto continue dissolving. Thus, the RSM-optimized conditions are economically superior and scientifically sound, prioritizing the co-recovery of both valuable metals.

Table summarizes the major element composition of the loaded leach solution obtained under the optimized conditions. The solution was predominantly enriched in Al (1.9 g/L) and K (1.8 g/L), alongside a minor Na content (0.39 g/L), which aligns with the selective dissolution of the feldspathic phases (microcline and albite). Several impurity elementsincluding Ca, Ba, Fe, Si, Mg, and Tiwere also detected at the 3.6–68 mg/L concentration range. Notably, the presence of 21 mg/L Co in the leachate is attributed to the abrasion of the milling media, as confirmed by the solid residue analysis (Table ), where the abraded WC remained undissolved, while the bulk of Co was leached and transferred into the solution.

8. Major Element Content of the Loaded Leach Solution Obtained under Optimum Conditions.

element	Al	K	Na	Ca	Ba	Fe	Si	Mg	Ti	Co
unit	g/L			mg/L
concentration	1.9 ± 0.1	1.8 ± 0.1	0.39 ± 0.3	68	31	26	22	7.5	3.6	21

9. Chemical Composition of the Solid Residue Obtained from the Leaching Process Conducted under Optimum Conditions .

main oxide	SiO2	Al2O3	Fe2O3	CaO	MgO	Na2O	K2O	TiO2	WO3	Co3O4	LOI
%	90.87	2.45	0.04	0.06	0.06	0.30	0.84	0.07	1.13	0.05	4.85

^aLOI: Loss on Ignition (1000 °C).

To verify the consistency of the experimental data, mass balance calculations were as performed for both aluminum and potassium based on the feed ore composition (Table ), the leach solution composition (Table ), and the solid residue composition (Table ). Based on the compositional data for the ore, the loaded leach solution, and the solid residue, the mass balance closure for aluminum and potassium was calculated. For aluminum, the input from 12.5 g of ore (16.33% Al₂O₃) corresponds to 1.0806 g Al, while the outputs are 0.950 g Al in the solution (1.9 g/L × 0.500 L) and 0.126 g Al in the residue (9.72 g residue × 2.45% Al₂O₃ × 54/102), giving a total output of 1.076 g Al and thus a closure of 99.6%. For potassium, the input from the ore (10.60% K₂O) is 1.0995 g K, with outputs of 0.900 g K in the solution (1.8 g/L × 0.500 L) and 0.0677 g K in the residue (9.72 g × 0.84% K₂O × 78/94), yielding a total output of 0.9677 g K and a closure of 88.0%. The nearly complete mass balance for aluminum (99.6%) validates the analytical procedure and confirms that the mass balance for this element is essentially closed. The lower closure for potassium (88.0%) indicates an apparent loss of approximately 12% of K, which can be attributed to some losses during washing and filtration, adsorption of K⁺ ions onto fine particles not recovered in the filtrate, the possible formation of soluble chloro-aluminate complexes (e.g., KAlCl₄) that may be incompletely measured by ICP-OES, and inherent analytical uncertainties.

3.4. Recovery of Al₂O₃ and KCl from the Leach Solution

As described in Section , the loaded leach solution obtained under the optimized leaching conditions was directly evaporated without prior pH adjustment or purification treatment. The evaporation step removed excess water and HCl (which was condensed and recovered for recycling), leading to the co-crystallization of hydrated AlCl₃ and KCl. The resulting mixed salt cake was then subjected to a two-stage roasting process to separate aluminum from alkali chlorides.

In the first stage, the salt mixture was roasted in a tube furnace at 1150 °C for 60 min. During this step, KCl volatilized, while alumina remained in the crucible as a solid oxide product. The volatile alkali chlorides were transported by the thermal gradient to the cooler zone of a borosilicate glass tube within the furnace, where they condensed and were collected as a crystalline deposit. The alumina product remaining in the crucible was subsequently re-roasted in a muffle furnace at 1200 °C and 1250 °C (each for 60 min) to examine the effect of temperature on product purity and crystalline phase transformation.

3.4.1. Characterization of the Alumina Product

The chemical compositions of the alumina products obtained at different roasting temperatures are presented in Table . The Al₂O₃ content increased from 88.8% at 1150 °C to approximately 92.3–92.6% at 1200 °C and 1250 °C. This improvement is directly correlated with the marked decrease in both moisture (at 300 °C) and loss on ignition (at 1000 °C) valuesfrom 3.69% at 1150 °C to ∼0.10% at higher temperatures. At 1150 °C, the 60 min roasting time is apparently insufficient to completely decompose intermediately hydrated and/or basic chloro-aluminate phases (e.g., residual Al­(OH)­Cl₂, AlOCl, or structurally bound hydroxyl groups in transitional alumina), which retain volatile components such as water and HCl. When the roasting temperature is raised to 1200 °C or 1250 °C, the additional thermal energy drives these decomposition reactions to completion, fully releasing the remaining H₂O, HCl, and trace chloride volatiles. The near-zero LOI values at these higher temperatures confirm that the product has become chemically stable and fully converted to nonhygroscopic α-Al₂O₃. This is consistent with the observed increase in α-Al₂O₃ content from 2.4% at 1150 °C to approximately 20–24% at 1200–1250 °C (Table ).

10. Chemical Composition of Alumina Products Obtained at Different Roasting Temperatures.

major oxide	roasting temperature (°C)
	1150	1200	1250
	(wt.%)
Al2O3	88.8	92.3	92.6
Fe2O3	0.665	1.02	0.922
SiO2	1.95	2.12	1.64
K2O	0.785	0.782	0.923
Na2O	0.349	0.04	0.175
CaO	1.87	1.62	1.71
TiO2	0.105	0.164	0.13
BaO	0.819	0.985	0.855
Co3O4	0.383	0.365	0.472
WO3	0.126	0.091	0.189
moisture (300 °C)	3.69	0.11	0.11
LOI (1000 °C)	3.69	0.1	0.09
α-alumina	2.4	23.9	20.8

The presence of Co₃O₄ and WO₃ in these products originates from the WC–Co milling media used in the mechanical activation step. As discussed in Section , this contamination is eliminated when sintered alumina (SA) milling media are employed, which also improves the Al₂O₃ product purity to over 94% and yields a white product instead of the blue-colored alumina obtained with WC–Co media. Nevertheless, the data in Table demonstrate that the evaporation–roasting route is effective in producing alumina with a purity suitable for many ceramic and refractory applications and that higher roasting temperatures (≥1200 °C) are beneficial for both volatile removal and α-Al₂O₃ crystallization.

3.4.2. Characterization of the Condensed Alkali Chloride Product

The salt product condensed on the cooler zones of the borosilicate tube during roasting at 1150 °C was analyzed by the XRD method. All diffraction peaks corresponded to KCl or NaCl (Figure ), with the most intense peaks belonging to KCl. This confirms that the volatilization–condensation process effectively separates the alkali chlorides from alumina.

11.

XRD pattern of the condensed alkali chloride product.

Quantitative analysis by FAAS revealed that the condensed salt contained 81.3% KCl and 17.3% NaCl. The presence of NaCl is attributed to the albite (NaAlSi₃O₈) component of the ore, which dissolves during HCl leaching and follows the same evaporation–roasting–volatilization–condensation pathway. The remaining minor components were below the detection limits of the analytical method. This KCl–NaCl mixture can be further purified to obtain high-purity KCl (>99%) by established techniques such as fractional crystallization or selective dissolution, which are the subject of ongoing studies.

As an alternative to the evaporation–roasting route, antisolvent crystallization using acetone has been demonstrated to recover KCl directly from acidic leach solutions with >83% K recovery and >99.9% purity.

3.4.3. Overall Recoveries and Mass Balance

Based on the loaded leach solution concentrations (Table ) and the final product masses, the overall recovery of K as condensed KCl was estimated at approximately 85–95% of K present in the loaded leach solution. The aluminum recovery as solid Al₂O₃ was essentially quantitative (>99%), with only minor mechanical losses during crucible transfer. The remaining K (5–15%) either remained in the crucible as a minor impurity in the alumina matrix or deposited as uncondensed fine particles in the furnace exhaust. These values are consistent with our previous experimental results and with similar chloride-roasting studies. ,

In summary, the evaporation–roasting scheme successfully demonstrates the feasibility of recovering both Al₂O₃ (as a solid product) and KCl (as a condensate product) from the loaded HCl leach solution, without the need for using additional reagents such as NaOH or lime. This integrated approach preserves the chloride circuit, enables acid recycling, and produces viable products, thereby supporting the zero-waste viewpoint of the POTASSIAL process.

3.5. Overall Process Evaluation and Mechanistic Insights

The experimental results presented in the preceding sections demonstrate that intensive mechanical activation of K-feldspar ore, followed by HCl leaching and subsequent evaporation–roasting, constitutes a viable route for the co-recovery of aluminum and potassium as Al₂O₃ and KCl, respectively. However, to fully appreciate the scientific and practical significance of the process, several key mechanistic, operational, and economic aspects require integrated discussion.

3.5.1. Mechanistic Interpretation of Amorphization and Leaching Behavior

The near-complete XRD-amorphization of microcline and albite after 120 min of intensive milling, contrasted with the partial XRD-amorphization of quartz (only 13.91%), reflects ultimate differences in crystal chemistry and bond topology. As noted in Section , feldspars contain weaker Al–O bonds and charge-balancing cations (K⁺, Na⁺) that introduce local lattice distortions, making the framework highly prone to mechanical collapse. In contrast, quartz’s homogeneous Si–O–Si network requires cooperative bond failure along specific crystallographic directions, imparting greater resistance to amorphization. , The mechanochemical bond breakage perspective aligns with recent conclusions on the reaction mechanisms of alkali-activated systems, where different amorphous aluminosilicate phases dissolve via distinct pathways.

FT-IR analysis (Figure ) provided direct spectroscopic evidence for these structural modifications. The activated sample exhibited significant changes in the framework vibrations: the Si–O–Si asymmetric stretching band near 1000 cm^–1 broadened and shifted slightly from approximately 1000 cm^–1 in the raw ore to 999.90 cm^–1 after milling, indicating disruption of the regular tetrahedral network and increased bond-length/bond-angle distortion. Concurrently, the band near 779 cm^–1, attributed to Si–O–Si symmetric stretching or Al–O vibrations, became more pronounced, reflecting the weakening and partial cleavage of Al–O bonds within the aluminosilicate framework. These spectral changes confirm that mechanical activation induces not only surface hydroxylation (as evidenced by the new O–H stretching band at ∼3400 cm^–1) but also significant lattice strain and bond rupture, particularly affecting Al–O–Si linkages. Farmer, Baláž, and Takacs have also reported that similar FT-IR changes are associated with structural disorder and amorphization in mechanically activated feldspar and other aluminosilicate mineral systems.

The enhanced leachability of the activated ore is directly attributable to this amorphization. The loss of long-range ordering weakens the aluminosilicate framework, reducing the activation energy required for proton attack. Moreover, the generation of fresh surfaces with coordinatively unsaturated Al–O and Si–O bonds creates highly reactive sites that readily undergo hydration and subsequent acid dissolution. The weakening of Al–O bonds, as evidenced by the changes in the 779 cm^–1 band, is particularly significant because these bonds are preferentially attacked during acid leaching, directly enhancing Al extraction. However, to fully interpret the dissolution behavior of the activated feldspar in HCl, it is necessary to consider the comprehensive mechanistic framework developed for aluminosilicate dissolution in acidic media.

The dissolution of aluminosilicates has been the subject of extensive theoretical debate, with several distinct mechanistic models proposed. The earliest and most influential of these is surface complexation models (SCMs), which postulate that the dissolution rate is controlled by the concentration of specific protonated surface species, particularly at mineral edge sites. Wieland and Stumm demonstrated that the rate-limiting step for kaolinite dissolution is the detachment of Al centers, with the activated complex exhibiting an Al/H⁺ stoichiometry that varies between 1:1 at edge surfaces and 1:3 at gibbsite surfaces. Building on this concept, Oelkers et al. improved the surface precursor complex model, disagreeing that the reaction rate is not solely a function of pH but is instead controlled by the decomposition of an Al-deficient, silica-rich surface precursor complex. This model elegantly explains why dissolution rates are often inversely dependent on aqueous Al concentration, thereby incorporating chemical affinity directly into the rate law. An alternative family of models focuses on the formation of leached layers, proposing that dissolution rates are governed by the diffusion of ions through a surface-altered, silica-enriched layer. In a related sequential mechanism, Oelkers et al. demonstrated that, under acidic conditions, tetrahedral Si–O bonds in muscovite structure break only after adjacent tetrahedral Al is removed via a proton–exchange reaction. This sequential pathway stands in direct contrast to the parallel-process theories proposed by other researchers. In this context, Crundwell presented a novel potential-coupled, parallel-process theory that departs significantly from both SCMs and leached-layer models. The key tenets of Crundwell’s theory are as follows: (i) The dissolution rate is controlled by the breaking of the metal–oxygen bond. Since the Al–O bond energy (≈3816 kJ mol^–1) is substantially weaker than the Si–O bond energy (≈13100 kJ mol^–1), , the aluminum–oxygen bonds in aluminosilicates are preferentially broken, while silicate tetrahedra (SiO₄ ^4–) leave the surface intact. (ii) Dissolution proceeds with two parallel, independent steps: cation (e.g., Al³⁺, Na⁺) and anion (SiO₄ ^4–) removal, coupled by the potential difference across the Helmholtz layer. For stoichiometric dissolution, the rates of cation and anion removal are equal, which allows for the elimination of the potential term from the rate expression. (iii) Crundwell highlights the critical observation by Casey et al. that the dissolution rate of plagioclase feldspars increases exponentially, with more substitution of aluminum for silicon in the tetrahedral structure. This strongly suggests that Al–O bond breaking directly controls the overall dissolution rate, reinforcing the treatment of Al³⁺ sites as the primary cationic centers.

Significantly, Crundwell analyzed earlier surface complexation and precursor models proposed for randomly assuming fractional stoichiometry to obtain half-order kinetics. In contrast, his theory derives the characteristic 0.5 reaction order naturally from the potential-dependent bond-breaking steps and the coupling of independent cation and anion fluxes. This represents a significant departure from traditional geochemical dissolution models and provides a unified explanation for the dissolution of not only aluminosilicates but also quartz, orthosilicates, and metal oxides.

Returning to experimental results, the observed decrease in Al recovery above ∼5 M HCl (reflected by the negative quadratic coefficient C² = −8.65 in eq ) can be interpreted within this multimodel framework. In the context of surface complexation models, the decline may reflect a change in surface speciation: at very low pH (high H⁺ activity), the surface becomes saturated with protons, and further acid addition does not increase the dissolution rate. In the leached-layer framework, the formation of a silica-rich passivating layer may be accelerated at extreme acidities, limiting proton access to the unreacted mineral core. From Crundwell’s perspective, the decrease is attributable to speciation shifts in the solution phase: as the HCl concentration increases, the activity of Cl^– rises, shifting the equilibrium from hydrated Al³⁺(aq) toward neutral AlCl₃(aq) and anionic AlCl₄ ^– complexes. , These uncharged or negatively charged species exhibit significantly lower solubility in the low-water-activity, high-ionic-strength medium, leading to their co-crystallization or “salting-out” from the solution phase. This effectively removes aluminum from the measured solution phase, reducing the apparent recovery.

It is important to emphasize that these mechanistic interpretations are not mutually exclusive; rather, they operate at different scales and may dominate under different conditions. The near-complete amorphization of the feldspar structure ensures that the mineral is far from equilibrium, so the surface precursor complex and Crundwell’s parallel-process model both predict rapid initial dissolution. As the reaction proceeds and the silica-rich leached layer develops, diffusion limitations become more significant, , consistent with the observed mixed kinetic regime. The sensitivity of Al recovery to HCl concentration, however, is best explained by the chloride speciation, which is a solution-phase thermodynamic limitation rather than a surface reaction phenomenon. By integrating these complementary perspectives, we provide a clear, noncontradictory framework for interpreting the experimental Al leaching results, ensuring that our discussion is both rigorous and internally consistent.

3.5.2. Contamination Management and Product Purity

The presence of W and Co in the activated ore (Table ) and in the final Al₂O₃ product (Table ) is a direct consequence of abrasion from WC–Co milling media. While this contamination does not invalidate the leaching results, it imposes constraints on product purity for high-end or advanced applications. The complementary tests using sintered alumina (SA) milling media demonstrate (Table ) that switching to alumina-based media keeps Co and W contamination below ICP-OES detection limits while simultaneously improving the Al₂O₃ purity from 92.6% (with WC–Co) to 94.1% (with SA) and producing a white instead of blue product.

11. Effect of Different Milling Media on the Composition of the Final Alumina Product ^,

alumina product	major oxide, wt %										color
	Al2O3	Fe2O3	SiO2	Na2O	CaO	K2O	TiO2	BaO	WO3	Co3O4
WC-media	92.6	0.92	1.64	1.18	1.71	0.92	0.13	0.86	0.19	0.47	blue
SA-media	94.1	0.46	1.29	0.28	1.90	0.44	0.16	0.84	n.d	n.d	white

^an.d. = not detected (below ICP-OES detection limit: Co < 0.01 wt %, W < 0.01 wt %).

^bK-feldspar activated with WC–Co bowl and balls.

^cK-feldspar activated with a sintered alumina bowl and balls.

For smelter-grade alumina, typical specifications require Al₂O₃ ≥ 98.0% with individual transition metals below 0.02%, while chemical-grade alumina demands even lower limits (<10 ppm for most impurities) (DADCO, 2026; FAO, 2026). , With SA media, the remaining impurities (Fe, Si, Ca, Na) are largely derived from the ore itself and can be further reduced by optimizing the leaching and washing conditions. Thus, the contamination issue is not an inherent limitation of the process but a media-selection problem that has been successfully overcome. The selective leaching behavior of these contaminants is consistent with the well-documented hydrometallurgical recycling of cemented tungsten carbide: the cobalt binder phase readily dissolves in acidic HCl solutions, while tungsten carbide remains largely intact in the solid residue. − This partitioning arises from the fundamentally different chemical stabilities of the two phases, and the results from this WC–Co recycling studies directly support our observation that Co is present in the leachate (Table ) while W is retained in the leach residue (Table ).

3.5.3. Basis for HCl Selection and Integrated Process Advantages

While H₂SO₄ is cheaper and widely used in industrial alumina production, its application in the present system would introduce severe operational complications. The calcium present in the ore (0.34% CaO) would precipitate as gypsum (CaSO₄·2H₂O) during H₂SO₄ leaching, causing scaling, acid loss, and entrainment of Al and K. More critically, the generation of sulfate salts would preclude the thermal volatilization–condensation separation that is central to our process. In the chloride system, AlCl₃·6H₂O decomposes to Al₂O₃ at 400–500 °C, while KCl and NaCl volatilize at >800 °C, enabling a clean, reagent-free separation. This chloride-based flowsheet also allows the recovery of HCl from the off-gases, which can be recycled to the leaching stage, reducing net acid consumption and operating costs. In contrast, conventional CaCl₂ roasting methods for KCl production generate large volumes of calcium aluminosilicate residues as solid waste and require water leaching to recover KCl, resulting in lower product purity and higher environmental impact. , The proposed integrated evaporation–roasting route avoids these drawbacks, offering a zero-waste, closed-loop alternative.

3.5.4. Purification of the Loaded Leach Solution

Feldspar ores are generally associated with gang minerals containing mainly Fe, Ti, Ca, or Mg elements. During HCl leaching, these elements dissolve alongside Al and K, entering the solution in ionic form. If these impurities are not sufficiently removed from the loaded leach solution prior to the evaporation–roasting step, they co-crystallize with the aluminum and potassium chlorides. Consequently, upon thermal conversion of AlCl₃ to Al₂O₃, the final alumina product exhibits high levels of Fe₂O₃, CaO, MgO, and TiO₂. Similar behavior has been observed in the acid leaching of natural pumice, where co-extracted Ca and Fe led to alumina products with high CaO and FeO contents. Therefore, purification of the loaded leach solution is essential for producing higher-purity potash and alumina.

However, the integrated HCl-leaching process described in this work requires that the chemical characteristics of the loaded leach solutionparticularly its anion composition and pHremain largely unchanged. This constraint arises from the following requirements: (i) Minimal neutralizationThe use of concentrated HCl for leaching means that any neutralization step would consume neutralizing reagents and increase operating costs. Moreover, neutralization can introduce additional impurities (e.g., Ca²⁺ from lime) or generate insoluble by-products, complicating subsequent filtration and purification stages. (ii) Maximum removal of impurities without compromising the chloride mediumEffective removal of multivalent cations (Fe³⁺, Ca²⁺, Mg²⁺, Co²⁺) is critical, as these species negatively affect the physical and chemical properties of the final products, especially the alumina. At the same time, the chloride medium must be preserved to enable the subsequent thermal separation of KCl and AlCl₃. (iii) Maintenance of solute concentrationsPurification steps should not significantly dilute the K⁺ and Al³⁺ concentrations, as subsequent recovery operations (particularly crystallization) require sufficiently high metal loadings to maintain efficiency. If impurity removal reduces the concentration of these target species, a concentration step may be necessary, which must be carefully integrated into the purification scheme. ,

3.5.5. Kinetic Limitations and Intensification Pathways

The optimal leaching time of 36 h, while scientifically justified for achieving maximum recovery, is economically challenging for industrial hydrometallurgy. However, the RSM model reveals that temperature (A) and its quadratic term (A²) exert the strongest influence on both Al and K recoveries (p < 0.0001), while the time factor (B) has a comparatively smaller effect. This suggests that the leaching process is not simply time-dependent but operates in a mixed kinetic regime: an initial chemically controlled stage (rapid dissolution of the amorphized framework), followed by a diffusion-controlled stage, where proton transport through the evolving silica-enriched leached layer becomes rate-limiting. , This interpretation is further supported by the high silica content (90.87% SiO₂, Table ) of the leach residue, which acts as a passive barrier that slows the reaction at longer times.

The formation of a silica-rich surface layer during aluminosilicate dissolution is a well-established phenomenon, often described by the interfacial dissolution–reprecipitation mechanism, where the leached layer develops a porous, interconnected structure. , In our system, this porous layer not only limits mass transfer (contributing to the long leaching time) but also provides a high-surface-area substrate that is subsequently exploited for silica purification and SiC synthesis.

On the basis of this understanding, several intensification strategies could be explored to reduce the leaching time to industrially acceptable levels: (i) two-stage acid addition, where an initial high-concentration acid charge dissolves the reactive fraction, followed by a second acid charge or temperature ramp to sustain the driving force; (ii) high-pressure leaching to increase the boiling point and accelerate reaction kinetics; or (iii) dual-frequency ultrasonic assistance to disrupt the silica barrier and enhance mass transfer.

Beyond these direct intensification strategies, recent developments in mineral processing and chemical engineering offer complementary insights that further strengthen the industrial viability of the process. The quantitative analysis of dewatering and filter-cake pore structure under ultrahigh pressures provides a mechanistic understanding of solid–liquid separation efficiency, which directly addresses a critical bottleneck in our flowsheet: the difficulty of filtering fine and highly activated particles after prolonged leaching. This work suggests that high-pressure dewatering could significantly improve filtration rates and cake washing efficiency, thereby reducing downstream processing time and energy consumption. The water conservation mining strategy and spatiotemporal evolution analysis offer a comprehensive framework for sustainable process water management. Their insights on closed-loop water circulation and real-time monitoring of water quality are directly applicable to our process, where HCl recovery and recycling are already integrated. Adopting such strategies would further minimize freshwater consumption and reduce liquid effluent generation, aligning the process with increasingly stringent environmental regulations. Finally, the synergy between biomass-driven reduction and acid leaching illustrates how renewable reductants can enhance the release of potassium and aluminum from aluminosilicate frameworks. While our current process does not employ a separate reduction step, this approach could be explored as a supplementary pretreatment to improve the reactivity of the feldspar lattice before acid leaching or as an alternative to fossil-based reductants in downstream carbothermal reduction steps. Taken together, these insights highlight that process intensification (via ultrasonic assistance and high-pressure dewatering), sustainable water management (via closed-loop strategies), and novel reductant pathways (via biomass) are critical levers for translating laboratory-scale results into industrial implementation.

3.5.6. Valorization of the Silica-Rich Residue

The leach residue, containing 90.87% SiO₂ (Table ), is not a waste stream but a valuable raw material. In this integrated process, the leach residue was successfully converted into high-purity amorphous and crystalline SiO₂ (97.7% purity) by mild alkali leaching and precipitation. This silica has then been used to synthesize calcium silicate (CaSiO₃) by reacting with marble industry waste (CaCO₃) and, more importantly, to produce silicon carbide (SiC) with a 98.9% purity via a novel carbothermal reduction route using biochar derived from almond shells, as more recently documented by Li et al. for similar biomass-assisted reduction processes. SiC is the primary economic driver of the overall process, as its market price (>$2500/ton) makes it significantly more valuable than KCl or Al₂O₃.

The pore structure evolution during acid leachingfrom the impenetrable feldspar framework to a highly porous, interconnected silica networkis not merely a passive outcome of the reaction but an active enabler of residue valorization. The enhanced accessibility to the pore surface facilitates subsequent alkali leaching and carbothermal reduction (for SiC synthesis), as demonstrated by our successful production of both products. This aligns with the broader concept of mechanochemical bond breakage, opening up new reactive pathways and the general principle that porosity development enhances fluid penetration and reactive surface exposure.

This valorization pathway transforms what would otherwise be a disposal problem into a revenue-generating product, aligning the process with the circular economy and zero-waste principles of the POTASSIAL project.

3.5.7. Techno-Economic and Environmental Outlook

A comprehensive techno-economic analysis of the entire POTASSIAL flowsheet (including mechanical activation, leaching, evaporation–roasting, SiO₂ purification, CaSiO₃ and SiC synthesis) was performed based on a 100 ton/day feed capacity. The total fixed capital investment was estimated at ∼76 million USD, with annual operating costs of ∼39.8 million USD. Under baseline market conditions, the process yields an internal rate of return (IRR) of 6.1% and a payback period of ∼12 years, which is marginal. However, sensitivity analysis reveals that the IRR increases to 12% and 22.8% with a 10% and 30% increase in SiC price, respectively, reducing the payback period to as low as 4 years. This indicates that the process economics are highly dependent on the SiC market value, which is expected to rise due to increasing demand in electronics, defense, and automotive sectors.

The life cycle assessment (LCA) using SimaPro 9.6 (TRACI 2.1 method) shows that while the POTASSIAL process has higher CO₂-equivalent emissions per product unit compared to conventional routes, its closed-loop operation and zero-waste generation offer substantial long-term environmental benefits: no saline effluents unlike brine potash extraction, no red mud unlike the Bayer alumina process, and reduced fossil carbon input owing to biochar use in SiC synthesis. Adopting closed-loop water management strategies, as discussed in the context of sustainable mining (Feng et al., 2025), would further minimize freshwater consumption. Thus, the process is environmentally superior despite its current cost challenges, and its financial viability can be significantly improved by optimizing product portfolios and capitalizing on the growing SiC market.

4. Conclusions

This study investigated the recovery of potassium and aluminum from mechanically activated K-feldspar ore via HCl leaching, followed by evaporation–roasting for the separation of Al₂O₃ and KCl. The following conclusions can be drawn from the experimental results:

• 1.Mechanical activation is necessary for rendering K-feldspar leachable. Intensive milling (ball-to-ore ratio of 30, 400 rpm, 120 min) caused complete amorphization (100%) of microcline and albite, while quartz remained partially crystalline (13.91%) due to its homogeneous Si–O–Si network, which resists structural collapse. The activated ore exhibited a sixfold increase in specific surface area from 4.68 to 28.05 m²/g and a significant increase in LOI from 0.46% to 2.88%, attributed to the formation of surface hydroxyls (Si–OH, Al–OH) and adsorbed water, as confirmed by FT-IR analysis. Under identical leaching conditions, the raw ore yielded only 2.83% Al extraction, whereas the optimally milled sample achieved an extraction efficiency of 87.72%, demonstrating the critical role of mechanical activation.
• 2.HCl concentration, temperature, and time are the principal leaching parameters. Response surface methodology (RSM) optimized the leaching conditions as 4.88 M HCl, 100 °C, and 36 h, yielding 82.08% Al and 84.20% K recoveries. The negative quadratic coefficient for HCl concentration (C² = −8.65) reflects the lower solubility of AlCl₃·6H₂O in concentrated HCl (>5 M) due to the formation of neutral and anionic chloro-aluminum complexes (AlCl₃(aq), AlCl₄ ^–) and reduced water activity. The kinetic analysis indicates a mixed regime: rapid chemical dissolution, followed by diffusion control through a silica-enriched leached layer, suggesting that process intensification (e.g., two-stage leaching, ultrasonic assistance, or high-pressure operation) could reduce the leaching time from 36 h to industrially acceptable durations. The mass balance analysis demonstrates excellent Al recovery with 99.6% closure, confirming the reliability of the experimental data. The K balance, however, shows 88.0% closurean ∼12% deficit likely due to washing losses, adsorption on fine particles, unmeasured chloro-aluminate complexes, and analytical uncertainty. Addressing these factors through optimized filtration and K speciation studies will be essential for improving the overall process efficiency.

• 3.HCl is the tactical lixiviant for process integration. Unlike H₂SO₄, which would precipitate CaSO₄ from the ore’s calcium content, HCl generates chloride salts that enable a clean thermal separation: during roasting at 1150–1250 °C, AlCl₃ converts to Al₂O₃, while KCl along with NaCl from the albite component volatilizes and condenses in the cooler zone of the furnace. The condensed salt mixture contains 81.3% KCl and 17.3% NaCl; however, the KCl fraction can be further purified to fertilizer grade or higher purity by established techniques (e.g., fractional crystallization or selective dissolution), which are the subject of ongoing studies. This reagent-free separation avoids the environmental burdens of conventional CaCl₂ roasting and pH-controlled precipitation.
• 4.Contamination from WC–Co milling media is fully manageable. The presence of W and Co in the final Al₂O₃ product is eliminated by switching to sintered alumina milling media, which also improves Al₂O₃ purity from 92.6% to 94.1% and prevents product coloration. This media substitution introduces no new impurities, as the wear debris is Al₂O₃ itself and does not compromise the amorphization degree or leaching efficiency.
• 5.The silica-rich leach residue is a valuable source. The siliceous residue of HCl leaching was successively converted into high-purity silica (97.7% SiO₂), CaSiO₃ from marble waste, and SiC (98.9% SiC) via biochar-driven carbothermal reduction. SiC is the primary economic driver of the process, and its market price critically determines overall profitability.
• 6.The integrated process is technically feasible but economically sensitive. The POTASSIAL process achieves zero-waste valorization of feldspathic ore, recovering multiple products (Al₂O₃, high-purity KCl, SiO₂, CaSiO₃, and SiC) without generating liquid or solid effluents. The techno-economic analysis shows that baseline economics (IRR ≈ 6.1%, payback ≈ 12 years) are marginal but become highly attractive (IRR > 22%, payback <4 years) with a 30% increase in SiC price. The LCA confirms long-term environmental benefits due to closed-loop operation, HCl recycling, and avoidance of red mud, saline brines, and fossil carbon in SiC synthesis.

In summary, the combined results demonstrate that the POTASSIAL approach, comprising mechanical activation, HCl leaching, evaporation–roasting, and residue valorization stages, successfully achieves near-complete recovery of K and Al from K-feldspar while generating zero solid waste. The chloride-based chemistry enables selective separation of KCl by volatilization, avoiding the environmental problems of conventional CaCl₂ roasting or pH-controlled precipitation. While the optimal leaching time and current product purity levels require further optimization, the process offers a technically feasible and environmentally sustainable alternative for the valorization of feldspathic ores, particularly in regions lacking evaporitic potash deposits.

The manuscript was written through contributions of all authors. All authors have given approval to the final version of the manuscript.

This work was supported by the institutions listed below. HORIZON EUROPE Widening Participation and Strengthening the European Research Area, ERA-MIN3 2021 Joint Call Secretariat, (POTASSIAL- Project No: 27). The Scientific and Technological Research Council of TurkeyTÜBİTAK (Projects Nos 122N039, 122N058, 122N059 and 122N060). İnönü University Scientific Research Projects Coordination Unit (Project Nos: FBG-2022-2978, FBA-2023-3131, FED-2024-3491 and FDK-2024-3710). Grant Agency of the Ministry of Education, Science, Research and Sport of the Slovak Republic (Project No: 2/0039/26). Slovak Research and Development Agency under Contract No. APVV-24-0353.

The authors declare no competing financial interest.