Study on Efficient Separation of Amorphous Silica from High-Alumina Coal Gangue

Abstract

Coal gangue, a major industrial solid waste from coal mining and processing, requires efficient alumina and silica separation for high-value utilization. This study focused on mineral reaction mechanisms and characteristics of coal gangue during calcination and alkaline leaching. Results showed calcination at 900–1200 °C altered its phase composition, affecting silica separation efficiency, with the optimal calcination range being 960–1120 °C. Poorly crystallized mullite and Al₂O₃ in calcined gangue were insoluble under low-alkaline and low-temperature conditions. On the contrary, amorphous silica is soluble and forms a sodium silicate solution in the proper alkaline conditions. This characteristic facilitates the efficient separation of alumina and silica. It was determined that the suitable conditions for silica removal from coal gangue are as follows: 1080 °C calcination for 90 min, leaching at 75 °C with 200 g/L NaOH (solid–liquid ratio of 1:4) for 4 h. Under these selected conditions, the silica leaching efficiency was 77.31%, the alumina leaching efficiency was 12.21%, the Na₂O content in the leached residue was 1.94%, and the mass ratio of alumina to silica (A/S) in the leached residue increased from 0.88 to 3.42. A potential desilication mechanism was also analyzed.

1. Introduction

Coal gangue is a substantial industrial solid waste generated in the process of coal mining and processing [1]. The main chemical composition of coal gangue is Al₂O₃ and SiO₂, and the key mineral composition of gangue is kaolin, illite, quartz, etc. [2]. The stockpile of coal gangue has grown to more than 7.7 billion tons [3] and is escalating annually by 300–350 million tons [4,5,6,7]. Most gangue is disposed by stockpile, which occupies a large amount of land and causes serious risks of environmental problems [8,9,10,11,12,13]. After years of efforts, although the comprehensive utilization rate of coal gangue in China has exceeded 70%, it is still mainly used for filling roadbeds and low-lying areas, road construction, simple reclamation, landfill, and other direct use, accounting for 35%–45% of the total amount of the gangue; gangue combustion in power plants and use of other fuels accounts for 10%–15%; gangue bricks and other building materials account for 5%–10%; and the filling of underground mining accounts for 3%–7% [14,15,16,17]. These modes of utilization are all low-value utilization.

The western regions of China (Ordos Basin) are rich in high-alumina coal resources, which is a large high-quality power coal base in China. The prospective reserves are more than 100 billion tons, the proven reserves are 58 billion tons, the mining output is over 200 million tons per year, and the associated high-alumina coal gangue (alumina content exceeds 35%) is about 20 million tons per year [18]. High-alumina coal gangue can be used for high-value utilization of alumina-based products based on its high alumina content and potential mineral resources [19].

The extraction of alumina, silicon, and other valuable elements from coal gangue is a crucial method for exploiting its high value. Various technical routes for alumina extraction from coal gangue have been reported, primarily categorized into an alkaline process and an acid process [20,21,22,23,24,25]. The acid method operates on the principle that alumina dissolves in the liquid phase with sulfuric acid or hydrochloric acid, while silicon remains in the solid phase, enabling alumina–silicon separation through solid–liquid separation. The acid process has the advantages of a high alumina extraction rate, short process flow, and fewer by-products. However, these processes have the disadvantages of generating large amounts of wastewater, the need for corrosive-resistant equipment, and high iron content in the intermediate products [26,27,28,29]. The alkali process is currently carried out in industrial applications and can be subdivided into a lime sinter process, soda-lime sinter process, pre-desilication and soda-lime sinter process, et al. [30,31,32,33]. The alkali process is characterized by technical maturity, minimal equipment demands, and direct applicability of products in aluminum electrolysis processes. Yet, challenges like lengthy process flows, substantial material flows, and increased by-product generation exist. In comparison to the acid process, the alkali method emerges as a more promising avenue for coal gangue resource utilization. Therefore, the difficulty of separating alumina and silica components hinders the further efficient utilization and high value of coal gangue.

Calcination of coal gangue to alter the phase composition and subsequent treatment of the calcined coal gangue by alkaline leaching is a potential way to separate alumina and silica in alumina-silica-bearing minerals [34,35,36]. Some studies have indicated that the inclusion of additives can enhance the activation effect during the calcination of coal gangue [37]. Li et al. [6,33] utilized hematite as an additive and implemented a reduction roasting–alkaline leaching process, achieving a silica leaching efficiency of approximately 95%. Song et al. [38]. demonstrated that calcination of coal gangue mixed with calcium hydroxide can significantly improve the activity of coal gangue. Guo et al. [39]. observed that with calcination temperatures ranging from 800 to 900 °C with Na₂CO₃ to coal gangue weight ratios of 0.8–1, alumina extractions exceeding 90% were obtained. When Lan et al. [40] employed KOH as an additive, with a coal gangue-to-additive ratio of 1:2, a microwave oven power of 595 W, and time of microwave activation of 20 min, a much higher extraction percent of Al (>50) was obtained. Previous research on these additives has primarily concentrated on alumina extraction from coal gangue. Nevertheless, incorporating additives into the calcination process presents challenges, such as the difficulty in controlling the calcined atmosphere, process complexity, and high alumina dissolution losses.

To address the treatment and resource utilization of coal gangue in coal-rich regions, there is an urgent necessity to develop and implement high-value utilization technologies for coal gangue, enabling large-scale reduction of coal gangue. The key to achieving high-value utilization of coal gangue lies in achieving efficient and cost-effective separation of alumina and silica. One of the key advantages of the calcination–alkaline leaching process is its utilization of self-carried carbon during calcination, reducing the reliance on external fuels. Enhanced desilication efficiency translates to higher resource utilization rates for coal gangue. Currently, there is a lack of research on the efficient separation of alumina and silica utilizing the intrinsic phase transformation properties of coal gangue itself, without the need for additives and employing a straightforward calcination treatment process. Current research outcomes in this field are predominantly derived from laboratory-based studies, and there remains a scarcity of practical cases for industrial application.

In this paper, the removal of silica from coal gangue through a calcination–alkaline leaching process was investigated. The mineral phase transformation of coal gangue was investigated during calcination ranging from 900 °C to 1200 °C. The characteristics of mineral phase transformation of the calcined clinker under alkaline leaching conditions were analyzed. Furthermore, the effects of desilication efficiency at different alkaline leaching temperatures was assessed. A potential desilication mechanism was speculated. This process results in the production of silica solution and high-alumina containing residue slag, both of which have practical applications. This study was experimentally validated on a 10,000-ton/year pilot platform, and the desilication performance achieved was consistent with the expected outcomes. The findings of this study carry considerable significance for the high-value utilization of coal gangue and the industrial translation of this technology—with particular relevance to low-grade bauxite ores rich in kaolinite (a typical high-silica aluminum resource). Furthermore, the microstructure of crystal transformation mechanism during calcination, the kinetics of the desilication reaction, and the high-value utilization of the desilication solution will be the core focuses of further research.

2. Materials and Methods

2.1. Materials

The analytical grade chemical reagent sodium hydroxide (NaOH) was procured from Sinopharm Chemical Reagent Co., Ltd., Tianjin, China and utilized for the alkaline leaching experiments. The coal gangue raw materials used in this experiment were gathered from Jinzhengtai coal mine in the Jungar coalfield, Inner Mongolia, China. The coal gangue (approximately 50 kg) that met the requirements was separated using X-ray intelligent separation technology (provided by Ganzhou Good Friend Co., Ltd., Ganzhou, China). After being dried in an oven at 105 °C for 12 h, the coal gangue was crushed and ground until its particle size was less than 150 μm. The chemical composition of the coal gangue is presented in

Table 1. Chemical composition in coal gangue and clinker (wt.%).

Sample	Al2O3	SiO2	CaO	Fe2O3	Na2O	TiO2	MgO	K2O	LOI
Coal gangue	37.28	42.72	0.05	0.12	0.11	0.62	0.10	0.13	18.18
Clinker	45.77	52.39	0.06	0.15	0.14	0.75	0.12	0.16	0.02

.

The XRD analysis of the mineral compositions indicated that kaolinite was the main phase in the coal gangue, along with a small quantity of boehmite (Figure 1).

2.2. Calcination Test

A 100.0 g sample of coal gangue was accurately weighed and gently placed in an 80 mm × 80 mm × 40 mm corundum crucible. Then, the loaded crucible was moved to a muffle furnace (KSL-1700X, Hefei Kejing Material Technology Co., Ltd., Hefei, China). The furnace temperature was gradually raised from ambient to specified levels (900–1200 °C in 20-degree intervals) at a rate of 5 °C/min. Upon the muffle furnace attaining the preset temperature, the sample was held at this temperature for a period of 90 min. After the calcination, the calcined coal gangue was left to cool naturally to room temperature. Then, it was taken out and used for the subsequent alkaline leaching experiments.

2.3. Alkaline Leaching Test

In the present study, alkaline leaching experiments were executed utilizing a temperature-regulated water bath pot. The optimization of these alkaline leaching processes was achieved by precisely modulating critical process parameters. In this research, the effects of alkaline leaching temperature on the alkaline leaching were systematically examined, the alkaline leaching temperature varied across the range of 45, 55, 65, 75, 85, and 95 °C, the alkaline leaching time was set to 4 h, the alkali concentration was specifically fixed at 200 g/L NaOH, and the solid-to-liquid ratio was determined as 15 g/60 mL. These parameters were selected based on preliminary exploratory experiments and industrial operational feasibility, ensuring sufficient desilication efficiency and compatibility with the production process. Subsequent to the leaching process, the solid residue was meticulously washed with deionized water maintained at a temperature of 90 °C. Subsequently, it was dried at a temperature of 105 °C for a duration of 3 h to facilitate further characterization. The entire experiment process can be represented in Figure 2.

The leaching efficiency of silica $(\mathsf{\eta }\left({\mathrm{S}\mathrm{i}\mathrm{O}}_2\right),\mathrm{\%})$ in calcined coal gangue is calculated using the following formula:

$$\eta \left({\mathrm{S}\mathrm{i}\mathrm{O}}_2\right)={\displaystyle \frac{\mathrm{M}1 {\mathsf{\phi }}_{\mathrm{a}}\left({\mathrm{S}\mathrm{i}\mathrm{O}}_2\right)-\mathrm{M}2 {\mathsf{\phi }}_{\mathrm{b}}\left({\mathrm{S}\mathrm{i}\mathrm{O}}_2\right)}{\mathrm{M}1 {\mathsf{\phi }}_{\mathrm{a}}\left({\mathrm{S}\mathrm{i}\mathrm{O}}_2\right)}}\times 100\%$$

(1)

where M1 is the mass of the calcined coal gangue(g), ${\mathsf{\phi }}_{\mathrm{a}}\left({\mathrm{S}\mathrm{i}\mathrm{O}}_2\right)$ is the silica content in the calcined coal gangue (%), M2 is the mass of the leached residue (g), and ${\mathsf{\phi }}_{\mathrm{b}}\left({\mathrm{S}\mathrm{i}\mathrm{O}}_2\right)$ is the silica content in the leached residue (%).

The leaching efficiency of alumina $(\mathsf{\eta }\left({{\mathrm{A}\mathrm{l}}_2\mathrm{O}}_3\right),\mathrm{\%})$ in calcined coal gangue is calculated using the following formula:

$$\eta \left({{\mathrm{A}\mathrm{l}}_2\mathrm{O}}_3\right)={\displaystyle \frac{\mathrm{M}1 {\mathsf{\phi }}_{\mathrm{a}}\left({{\mathrm{A}\mathrm{l}}_2\mathrm{O}}_3\right)-\mathrm{M}2 {\mathsf{\phi }}_{\mathrm{b}}\left({{\mathrm{A}\mathrm{l}}_2\mathrm{O}}_3\right)}{\mathrm{M}1 {\mathsf{\phi }}_{\mathrm{a}}\left({{\mathrm{A}\mathrm{l}}_2\mathrm{O}}_3\right)}}\times 100\%$$

(2)

where M1 is the mass of the calcined coal gangue(g), ${\mathsf{\phi }}_{\mathrm{a}}\left({{\mathrm{A}\mathrm{l}}_2\mathrm{O}}_3\right)$ is the alumina content in the calcined coal gangue (%), M2 is the mass of the leached residue (g), and ${\mathsf{\phi }}_{\mathrm{b}}\left({{\mathrm{A}\mathrm{l}}_2\mathrm{O}}_3\right)$ is the alumina content in the leached residue (%).

2.4. Thermodynamic Analysis

Thermodynamic analysis of the reaction equations was conducted using HSC 9.0 software, leveraging its built-in H, S, and Cp Thermochemical Database and Minerals Database.

To validate the reliability of these results, a manual cross-comparison was performed against thermodynamic data reported in the literature [41]. For the thermodynamic calculations, the properties of SiO₂ and Al₂O₃ were represented using the database data for quartz and γ-Al₂O₃, respectively. For kaolinite, its Gibbs free energy of formation (∆G_f) at 700–1600 K was derived via linear fitting as ∆G_f = 1.07T − 4119.53 with a correlation coefficient R² = 1.00. This expression was obtained by extrapolating the linear fitting results of kaolinite’s thermodynamic data (at 298.15–700 K) to the 700–1600 K temperature range [34].

2.5. Characterization

The phase and crystalline structures of materials were determined through X-ray diffraction (XRD, D8 Advanced, Bruker, Karlsruhe, Germany) analysis using Cu Kα radiation (λ = 0.1542 nm) and the 2theta test range of 5–80°. The morphology of samples was examined utilizing a Scanning electron microscopy (SEM, Hitachi S-4800, Tokyo, Japan) equipped with energy-dispersive spectroscopy (EDS, Oxford X-MAX20, High Wycombe, UK) detectors for compositional analysis. Chemical compositions of the calcined coal gangue and the leached residue were analyzed using X-ray fluorescence spectrometry (XRF, ARL PERFORM’X-4200, Thermo Fisher, Tewksbury, MA, USA). The chemical bonds in the calcined coal gangue and the leached residue were identified using a Fourier transform infrared spectrometer (FTIR, Nicolet iS5, Thermo Fisher, Madison, WI, USA) with spectral range of 400–4000 cm⁻¹. Thermogravimetric (TG) analysis was carried out using NETZSCH STA 449 C (NETZSCH-Gerätebau GmbH, Selb, Germany) from room temperature to 1300 °C at a heating rate of 10 °C min⁻¹ under an air atmosphere.

3. Results

3.1. Thermal Behavior and Thermodynamic Analysis of High-Alumina Coal Gangue During Calcination

To understand the thermal behavior of high-alumina coal gangue (a key step for optimizing its calcination and desilication processes), TG-DSC analysis was performed, with the curves presented in Figure 3. The TG curve reveals a significant mass loss (17.7%) in the temperature range of 400–700 °C, which corresponds to the dehydroxylation of kaolinite (the dominant aluminosilicate mineral in coal gangue). Furthermore, the initiation temperature of dehydroxylation for kaolin was ascertained to be 442.7 °C. An exothermic peak was detected at 993.06 °C, which serves as an indication of the transition temperature from kaolin to the spinel or mullite phase [33].

To explain the thermal behavior observed in TG-DSC curves, thermodynamic calculations were performed. High-alumina coal gangue is predominantly composed of two primary minerals: kaolinite and boehmite. With increasing calcination temperatures, the primary potential reactions are summarized as follows:

$$2\mathsf{\gamma }-\mathrm{AlOOH} = \mathsf{\gamma }-\mathrm{A}{\mathrm{L}}_2{\mathrm{O}}_3 + {\mathrm{H}}_2\mathrm{O}$$

(3)

Al₂O₃·2SiO₂·2H₂O (kaolinite) = Al₂O₃·2SiO₂ (metakaolin) + 2H₂O

(4)

3(Al₂O₃·2SiO₂) = 3Al₂O₃·2SiO₂ + 4SiO₂

(5)

$$4({\mathrm{Al}}_2{\mathrm{O}}_3·2{\mathrm{SiO}}_2)=3{\mathrm{Al}}_2{\mathrm{O}}_3·2{\mathrm{SiO}}_2+6{\mathrm{SiO}}_2+\mathsf{\gamma }-{\mathrm{Al}}_2{\mathrm{O}}_3$$

(6)

$$2{\mathrm{SiO}}_2+3\mathsf{\gamma }-{\mathrm{Al}}_2{\mathrm{O}}_3=3{\mathrm{Al}}_2{\mathrm{O}}_3·2{\mathrm{SiO}}_2$$

(7)

Thermodynamic analysis of the above reaction equations was performed using HSC 9.0 software, with results presented in Figure 4. The manually derived results show good consistency with the thermodynamic calculations from HSC 9.0. Under standard atmospheric pressure, the standard molar Gibbs free energy change (ΔrG^θ) of reaction Formula (3) remains positive at temperatures ranging from 0 to 700 °C, while that of reaction Formula (4) is positive between 0 and 400 °C. This indicates these reactions are non-spontaneous under the specified conditions—meaning they do not proceed spontaneously in the forward direction and require external energy input to proceed within these temperature ranges. Above 700 °C, kaolinite decomposes into metakaolin and H₂O, while boehmite concurrently breaks down into Al₂O₃ and H₂O. At temperatures exceeding 980 °C, metakaolin further decomposes into amorphous silica and mullite [6]. As illustrated in Figure 4, the conversion of metakaolin to mullite and amorphous silica proceeds readily during calcination, as reaction Formulas (4) and (5) exhibit negative ΔrG^θ values. With continued increases in calcination temperature, Al₂O₃ and amorphous silica can react to form mullite due to the negative ΔrG^θ of reaction Formula (7).

3.2. Separation of Amorphous Silica

3.2.1. The Impact of Calcination Temperature on Desilication Efficiency

To understand more comprehensively the leaching behavior of amorphous silica from gangue clinker at various calcination temperatures in an alkaline environment, the gangue clinker was subjected to alkaline leaching. This was carried out in a 200 g/L NaOH solution with a solid/liquid ratio of 15 g/60 mL at different alkali-desilication temperatures for 4 h.

As shown in Figure 5(a1,a2), alkali desilication of calcined gangue clinker at 55 °C revealed that the A/S ratio in leached residues increased from 0.87 to 1.87 with calcination temperatures rising from 900 °C to 1080 °C, then decreased to 1.17 at 1200 °C. Concurrently, the Na₂O content dropped from a peak of 4.83% to 0.22%. Silica leaching efficiency rose gradually from 8.22% to 56.57% within 900–1080 °C, but declined to 25.05% at 1200 °C. Alumina leaching efficiency remained below 10.0%, decreasing with higher temperatures.

Figure 5(b1,b2) depicts results at 65 °C: the A/S ratio surged from 0.86 to 2.67 between 900 and 1080 °C, then fell to 1.24 at 1200 °C. Na₂O content dropped from 8.45% to 0.18% across the same temperature range. Silica leaching efficiency peaked at 70.36% at 1080 °C (up from 5.50% at 900 °C), then declined to 31.29% at 1200 °C. Alumina leaching efficiency fluctuated between 5.0% and 10.0%.

At 75 °C (Figure 5(c1,c2)), the A/S ratio sharply increased to 3.42 at 1080 °C (from 0.85 at 900 °C), then decreased to 1.52 at 1200 °C. Na₂O content dropped from 10.55% to 0.34%. Silica leaching first rose steeply to 68.74% at 960 °C, then plateaued to a peak of 77.30% at 1080 °C, before falling to 44.61% at 1200 °C. Alumina leaching efficiency ranged from 5.0% to 15.0%.

As shown in Figure 5(d1,d2), alkali desilication at 85 °C showed that the A/S ratio in leached residues surged from 0.92 to 3.18 with calcination temperatures rising from 900 °C to 1060 °C, then gradually declined to 1.76 at 1200 °C. Na₂O content dropped from a peak of 12.36% at 900 °C to 1.17% at 1200 °C. Silica leaching efficiency spiked from 8.54% to 64.30% between 900 and 960 °C, rose to a peak of 75.22% at 1060 °C, then fell to 51.62% at 1200 °C. Alumina leaching efficiency fluctuated between 5.0% and 12.0%.

For the 95 °C condition (Figure 5(e1,e2)), the A/S ratio slowly increased from 0.85 to 2.44 as calcination temperature rose from 900 °C to 1140 °C, then declined to 1.82 at 1200 °C. Na₂O content decreased from 15.87% at 900 °C to 1.25% at 1200 °C. Silica leaching efficiency surged from 6.97% to 46.49% between 900 and 960 °C, climbed to a peak of 65.12% at 1140 °C, and dropped to 53.37% at 1200 °C. Alumina leaching efficiency remained in the 5.0%–13.0% range across all temperatures.

Table 2. Chemical composition of leached residues after desilication under different calcination temperatures (wt.%).

Calcination Temperature	Al2O3	SiO2	CaO	Fe2O3	Na2O	TiO2	MgO	K2O	LOI
900 °C	33.83	40.34	0.29	0.10	10.54	0.53	0.08	0.11	14.32
960 °C	60.22	24.84	0.25	0.24	4.28	0.99	0.14	0.18	6.89
1020 °C	67.14	21.33	0.36	0.44	2.90	1.07	0.16	0.19	5.78
1080 °C	70.63	20.64	0.45	0.45	1.94	1.16	0.18	0.20	3.13
1140 °C	68.79	24.11	0.55	0.49	1.56	1.05	0.16	0.18	2.26
1200 °C	57.87	38.25	0.29	0.22	0.31	0.91	0.14	0.18	1.22

Note: Leaching conditions: 75 °C, 4.0 h, NaOH concentration of 200 g/L, solid/liquid of 15 g/60 mL.

displays the chemical composition of leached residues post-desilication, illustrating how varying calcination temperatures affect the residue’s elemental distribution.

As indicated in Figure 5 and Table 2, based on the analysis of the A/S ratio, Na₂O content, silica leaching efficiency, and alumina leaching efficiency, the clinker exhibits an underburned state at calcination temperatures of 900–960 °C, reaches the optimal normal-burned state at 960–1120 °C, and enters an overcooked state when the temperature increases to 1120–1200 °C. Analysis results indicate that within the desilication temperature range of 55–75 °C, the silica leaching efficiency increases progressively with increasing temperature. However, when the temperature exceeds 85 °C, this ratio decreases progressively as the temperature continues to rise.

3.2.2. XRD Analysis

Comparative XRD analysis of the before and after resultant alkaline leached residue of calcined coal gangue is show in Figure 6.

The XRD diffraction pattern reveals four phases: mullite, Al₂O₃, anatase, and amorphous silica. Al₂O₃ and amorphous silica exhibit stronger diffraction. After alkaline leaching, there are three phases on the leached residue: Al₂O₃, mullite, anatase; the complete disappearance of the diffraction peak of amorphous silica at 20–22° (2θ) indicates that amorphous silica may have been dissolved in the alkaline solution. Weak diffraction peaks at 2θ = 45.86° and 2θ = 67.03° corresponding to the [400] and [440] lattice plane of Al₂O₃, respectively, are observed in the before and after alkaline leached residue. The amorphous silica is soluble in the alkaline solution, combined with the presence of Al₂O₃ in the alkaline leached residue, and represents the primary factor contributing to the increase in the A/S ratio within the residue.

3.2.3. FT-IR Analysis

Figure 7 displays the FTIR spectra of coal gangue, clinker, and leached residue, while

Table 3. FTIR characteristic band assignments for coal gangue, clinker, and leached residue.

Coal Gangue Band (cm−1)	Clinker Band (cm−1)	Leached Residue Band (cm−1)	Bond Type
3689/3650/3619	-	-	OH−
1116	1078	1109	Si-O (Si-O-Si/Si-O-Al)
1030/1009	-	-	Si-O-Si/Si-O-Al
-	-	967	Si-O-Al
939/913	-	-	Al-O-H
738/750/680/645	802/726	826/724	Si-O-Si/Si-O-Al
537/461/429	552/451/418	535/420	Si-O/Al-O

summarizes the characteristic absorption bands corresponding to these three samples.

For coal gangue, peaks at 3689/3650/3619 cm⁻¹ (kaolinite O-H stretches [42,43]), 1116/1030/1009 cm⁻¹ (Si-O/Si-O-Al stretches [44]), 939/913 cm⁻¹ (Al-O-H swing [45]), and 738/750/680/645 cm⁻¹ (Si-O-Si/Si-O-Al bending/deformation [10,45,46,47,48,49,50,51,52]) reflect its kaolinite-dominated structure.

For the calcined clinker, the disappearance of O-H-related peaks (3689/3650/3619 cm⁻¹) indicates hydroxyl removal via calcination; the broad peak at 1078 cm⁻¹ (assigned to Si-O (Si-O-Si/Si-O-Al) stretch vibration [44]), combined with the absence of sharp crystalline SiO₂diffraction peaks in XRD, suggests the presence of amorphous silica; 802/726 cm⁻¹ (Si-O-Si/Si-O-Al bending [46,47]) and a552/451/418 cm⁻¹ (Si-O/Al-O deformation [48,49]) correspond to the calcined silicoaluminate framework.

For leached residue, a further shifted broad Si-O peak (1109 cm⁻¹ [44]) and a distinct Si-O-Al peak (967 cm⁻¹ [48]) (plus 826/724 cm⁻¹ [46,47] and 535/420 cm⁻¹ [49,50]) reveal structural modifications from desilication, consistent with the change in Si-containing phases.

3.2.4. SEM-EDS Analysis

To further elucidate the desilication reaction characteristics of coal gangue clinker calcined and the corresponding alkaline leached residue, they were subjected to detailed SEM-EDS characterization, as depicted in Figure 8.

The energy-dispersive spectroscopy (EDS) mapping of clinker and alkaline leached residue is presented in Figure 8a and Figure 8b, respectively. The elemental C shown in Figure 8 is mainly due to the situation where the sample was put on a conductive adhesive that contains C. Within the labeled particle range, energy spectrum analyses revealed that the coal gangue contained 8.29 wt.% Al and 6.77 wt.%Si, as shown in Figure 8a. Following alkaline leaching, the residue contained 13.37 wt.% Al and 2.85 wt.%Si, as illustrated in Figure 8b. This outcome implies that a significant quantity of SiO₂ dissolves in alkaline solutions, resulting in a mass reduction and significant shifts in the relative proportions of Al and Si in the calcined gangue clinker as well as the alkaline leached residue. The major elements Al, Si, and O were uniformly dispersed within the particles not alkaline leached, with their respective contents closely aligned. In contrast, the distribution of the major elements Al and Si in the alkaline leached residue indicated a higher concentration of Al and a lower concentration of Si, while the concentration of element O remained almost constant. Upon comparing the elemental concentrations of calcined gangue clinker, a decrease in the concentration of elemental Si and a lesser decrease in the concentration of elemental Al were shown in the solid phase of alkaline leached residue. The elemental distribution patterns revealed by EDS mapping are consistent with the compositional profiles derived from energy spectrum analysis.

3.3. Mineral Phase Transformation of Coal Gangue During Calcination and Alkaline Leaching

In order to obtain more profound understanding of the phase transformation behavior of coal gangue under different calcination temperatures and the phase transformations happening during the alkaline leaching procedure, the XRD patterns of the coal gangue calcined for 90 min at various temperatures and the XRD analysis results of the leached residues are presented in Figure 9, with the leaching conditions being 75 °C, 4 h, a NaOH concentration of 200 g/L, and a solid/liquid ratio of 15 g/60 mL.

As shown in Figure 9(a1), when the calcination temperature increases from 900 °C to 920 °C, the disappearance of boehmite’s characteristic diffraction peaks indicates boehmite’s complete transformation into Al₂O₃ via Formula (3), while the vanishing of kaolinite’s peaks suggests its conversion to amorphous meta-kaolinite through Formula (4). During the decomposition of kaolinite and boehmite into an amorphous phase within the temperature range of 900–920 °C, a weak diffraction peak remains at 2θ = 25.303°; this peak corresponds to the [101] lattice plane of anatase. The existence of the anatase phase might be attributed to the approximately 1% TiO₂ contained in the coal gangue. Moreover, a broad peak, which represents the typical amorphous silica phase, is observed at 2θ = 20–24°. When the calcination temperature ascends from 940 °C to 960 °C, four phases can be detected in the XRD diffraction pattern: mullite, amorphous silica, Al₂O₃, and anatase. Weak diffraction peaks appear at 2θ = 16.43° and 2θ = 26.27°, corresponding to the [110] and [210] lattice planes of mullite, respectively.

In the leached residues of clinkers calcined at temperatures between 900 °C and 920 °C, zeolite A along with a small quantity of sodalite were detected in Figure 9(a2). The diffraction intensity of sodalite increased steadily, whereas the diffraction intensity of zeolite A declined gradually. The existence of zeolite A and sodalite within the leached residues shows that meta-kaolinite can engage in a reaction with the alkaline solution as per Formula (8) [32], causing the production of sodium aluminate solution and sodium silicate solution. In the following steps, Formulas (11) and (12) between sodium aluminate and sodium silicate solutions with distinct molecular ratios result in the formation of zeolite A and sodalite respectively [31]. These zeolite-like substances significantly increase the alkali content within the leached residue.

As depicted in Figure 9(b1,c1), when the calcination temperature ascends from 980 °C to 1120 °C, four phases can be detected in the XRD diffraction pattern: mullite, amorphous silica, Al₂O₃, and anatase. Weak diffraction peaks appear at 2θ = 16.43° and 2θ = 26.27°, corresponding to the [110] and [210] lattice planes of mullite, respectively. At the calcination temperatures of 940 °C to 1060 °C, the characteristic diffraction peaks of mullite are relatively weak. In comparison, the mullite peaks at 1120 °C are stronger than those at 940 °C. Based on previous research [31], amorphous meta-kaolinite might decompose into mullite and amorphous silica via Formula (5). Diffraction peaks are present at 2θ = 45.86° and 2θ = 67.03°, which, respectively, correspond to the [400] and [440] lattice planes of Al₂O₃ (PDF#10-0425). The characteristic diffraction peaks of Al₂O₃ that are detected are relatively conspicuous at calcination temperatures of 940 °C and 1120 °C. The emergence of the Al₂O₃ diffraction peaks can be ascribed to the transformation of boehmite at elevated temperatures via Formula (3) and the decomposition of kaolinite through Formula (6). As a result of the decomposition of metakaolin, there is a relative increase in the diffraction intensity of amorphous silica at the main peak position corresponding to 2θ in the 20–24° interval. The abundant amorphous silica in the products of coal gangue calcination at 940 °C and 1120 °C provides the foundation for efficient alumina–silica separation.

As depicted in Figure 9(b2,c2), the diffraction intensity of amorphous silica at the main peak position where 2θ is between 20 and 22° disappears because amorphous silica is soluble in the alkaline solution. Mullite, anatase, Al₂O₃, and a small quantity of Na₆Al₆Si₁₀O₃₂ are detected in the leached residues of clinkers that have been calcined at temperatures ranging from 940 °C to 1120 °C. As the calcination temperature of coal gangue increases to 980 °C and 1120 °C, the reaction between amorphous silica and sodium hydroxide generates sodium silicate through Formula (9), and active alumina and sodium hydroxide generate sodium aluminate through Formula (10) [35]. The reactions described by these equations explain the increase in the A/S ratio of the leached residue and the observed alumina leaching behavior, as shown in Figure 5. The existence of Na₆Al₆Si₁₀O₃₂ within the leached residues shows that amorphous silica and active Al₂O₃ can engage in a reaction with the alkaline solution through Formulas (9) and (10), causing the production of sodium aluminate solution and sodium silicate solution. In the following steps, Formula (13) between sodium aluminate and sodium silicate solutions with distinct molecular ratios results in the formation of Na₆Al₆Si₁₀O₃₂. Mullite and Al₂O₃ are present in the leached residue, which implies that, within the current reaction circumstances, the reactivity of these components with the alkaline solution is somewhat limited.

6NaOH + Al₂O₃·2SiO₂ = Na₂O·Al₂O₃ + 2Na₂SiO₃ + 3H₂O

(8)

2NaOH + SiO₂ (amorphous) = Na₂SiO₃ + H₂O

(9)

2NaOH + Al₂O₃ (active) = Na₂O·Al₂O₃ + H₂O

(10)

Na₂O·Al₂O₃ + 2Na₂SiO₃+ 6.5H₂O = Na₂O·Al₂O₃·2SiO₂·4.5H₂O + 4NaOH

(11)

Na₂O·Al₂O₃ + 6Na₂SiO₃+ 6H₂O = Na₂O·Al₂O₃·6SiO₂ +12NaOH

(12)

3Na₂O·Al₂O₃ + 10Na₂SiO₃ + 10H₂O = Na₆Al₆Si₁₀O₃₂ + 20NaOH

(13)

3Al₂O₃·2SiO₂ + 10NaOH = 6NaAlO₂ + 2Na₂SiO₃ + 5H₂O

(14)

3Na₂O·Al₂O₃ + 10Na₂SiO₃+ 22H₂O = Na₆Al₆Si₁₀O₃₂·12H₂O +20NaOH

(15)

When the temperature of calcination increases from 1140 °C to 1200 °C, as shown in Figure 9(d1), the characteristic diffraction peaks of mullite are highly conspicuous. This shows that with an increase in temperature, the crystallization form of mullite is more distinct. The occurrence of mullite diffraction peaks is due not only to the breakdown of metakaolin through Formulas (5) or (6) but also to the reaction of Al₂O₃ and amorphous silica in accordance with Formula (7). When the calcination temperature of coal gangue is increased to 1200 °C, the characteristic diffraction peaks of Al₂O₃ vanish. The reason could be that Al₂O₃ reacts with amorphous silica to form mullite via Formula (7). As the calcination temperature rises, the diffraction intensity of amorphous silica in coal gangue samples declines significantly.

In the leached residues of clinkers calcined at temperatures from 1140 °C to 1200 °C, in addition to the pre-existing mullite, a small quantity of Al₂O₃ was detected in Figure 9(d2). When the diffraction intensity of amorphous silica at the main peak position where 2θ is between 20 and 22°disappeared, no novel substances were detected. These findings imply that mullite and Al₂O₃ exhibit low reactivity towards the alkaline solution. The XRD analysis results are consistent with the leaching experiments depicted in Figure 5(c1,c2).

To gain deeper insights into the phase composition of alkaline leached residues under different desilication temperatures (55 °C, 65 °C, 75 °C, 85 °C, 95 °C), their XRD patterns are shown in Figure 10. When the desilication temperature increases from 55 °C to 75 °C, the same three phases—mullite, Al₂O₃, and anatase—are consistently detected in the XRD patterns of the leached residue. At 55 °C, the XRD pattern exhibits a broad bun peak (2θ = 20~25°) characteristic of amorphous SiO₂, suggesting incomplete dissolution of amorphous SiO₂ in the NaOH solution due to an insufficient desilication temperature. As the temperature increases to 65 °C, the integral intensity of the amorphous SiO₂ bun peak decreases, reflecting a reduced proportion of undissolved amorphous SiO₂ in the residue. This indicates that an elevated temperature enhances the solubility of amorphous SiO₂ in the NaOH solution, promoting its dissolution. At 75 °C, the characteristic bun peak of amorphous SiO₂ nearly vanishes, confirming complete dissolution of amorphous SiO₂ in the NaOH solution under sufficient desilication temperatures. At 85 °C and 95 °C, the XRD patterns of the leached residues reveal the emergence of zeolite P1 (Na₆Al₆Si₁₀O₃₂·12H₂O) and anatase, with the diffraction intensity of zeolite P1 being notably higher at 95 °C than at 85 °C. The presence of zeolite P1 and the disappearance of primary crystalline phases (mullite and Al₂O₃) indicate that amorphous silica and active Al₂O₃ react with the alkaline solution via Formulas (9) and (10), dissolving into sodium aluminate (NaAlO₂) and sodium silicate (Na₂SiO₃) solutions; and primary mullite decomposes via reaction (14), contributing additional aluminate and silicate ions to the solution. Subsequently, the molecular ratio-dependent Formula (15) between sodium aluminate and sodium silicate solutions triggers the crystallization of zeolite P1, as depicted by Formula (15).

To better characterize the chemical composition of alkaline leached residues under different desilication temperatures, their constituent profiles are presented in

Table 4. Chemical composition of leached residues after desilication under different desilication temperatures (wt.%).

Desilication Temperature	Al2O3	SiO2	CaO	Fe2O3	Na2O	TiO2	MgO	K2O	LOI
55 °C	61.47	33.56	0.50	0.23	0.47	0.99	0.14	0.18	1.02
65 °C	68.02	25.52	0.58	0.30	1.11	1.10	0.15	0.19	1.87
75 °C	70.63	20.64	0.45	0.45	1.94	1.16	0.18	0.20	3.13
85 °C	67.11	21.74	0.54	0.25	3.46	1.05	0.15	0.18	4.87
95 °C	54.98	26.43	0.33	0.24	8.11	1.02	0.14	0.17	8.89

Note: Calcination temperatures of 1080 °C, leaching conditions as follows: 4.0 h, NaOH concentration of 200 g/L, solid/liquid ratio of 15 g/60 mL.

. As the desilication temperature increases from 55 °C to 95 °C, the Na₂O content in the residues rises significantly; this trend can be ascribed to the formation of diverse zeolite phases. Importantly, these chemical analysis results exhibit a strong correlation with the XRD results shown in Figure 10.

In an effort to elucidate why desilication efficiency declines as desilication temperature increases, XRD characterization was performed on the alkaline leached residue (obtained at 95 °C) across different calcination temperature regimes, as shown in Figure 11. Below 960 °C, zeolite A and Faujasite-Na react rapidly under the studied experimental conditions. For calcination temperatures ranging from 960 to 1120 °C, primary mullite and amorphous SiO₂ undergo reactions with the alkaline solution according to Formulas (9) and (14), producing sodium silicate and sodium aluminate as intermediate products. These intermediates subsequently undergo a secondary reaction to generate zeolite P1. When the calcination temperature is elevated to 1140–1200 °C, the final solid product is a mixture of zeolite P1 and residual primary mullite. By comparing with Figure 9, it can be concluded that the primary mullite phase disappears within the calcination temperature range of 940–1100 °C. This indicates that at a desilication temperature of 75 °C, the alkaline solution only leaches amorphous SiO₂ without reacting with the primary mullite phase. However, when the desilication temperature is elevated to 95 °C, the reaction pathway changes: the alkaline solution first reacts with amorphous SiO₂, and after the majority of amorphous silica is consumed, it further reacts with the primary mullite phase. The decomposition of mullite leads to a significant increase in the liquid-phase Al₂O₃ concentration, which drives the in situ formation of a zeolite phase. This zeolite phase precipitates back into the solid residue, thereby inhibiting the leaching of silica and reducing its leaching rate.

3.4. Microscopic Morphological Changes of the Leached Residues After Alkaline Leaching

To obtain a more profound understanding of the microscopic morphological alterations of alkaline solution leached residues at various calcination temperatures, SEM analyses were carried out, as depicted in Figure 12. The leaching conditions are 75 °C, 4 h, a NaOH concentration of 200 g/L, and a solid/liquid ratio of 15 g/60 mL.

In Figure 12a, a multitude of zeolite A cubic particles, measuring only 1.0–3.0 μm, alongside small spherical sodalite particles measuring approximately 0.5–1.0 μm at 900 °C is shown. As the calcination temperature increases to 920 °C in Figure 12b, the amount of zeolite A cubic particles decreases, measuring only 0.5–2.0 μm, whereas the amount of sodalite particles increases, measuring approximately 1.0–2.0 μm. These SEM images align closely with the XRD analysis findings in Figure 6. As the calcination temperature increases to 940 °C in Figure 12c, a very small amount of sodalite particles can be observed and most of the particles are clustered flakes. As the calcination temperature increases from 960 °C to 1120 °C in Figure 12d–l, the leached residue exhibits a granular and aggregated state, with minor changes compared to the calcination clinker, and some large particles are present as calcination temperature increases. After leaching with an alkaline solution, the leached residue shows minimal structural changes.

As the calcination temperature progresses (Figure 12m–p), the leached residue exhibits a granular and aggregated state. Aggregation between particles is denser in this temperature range than that in the lower temperature range (960–1120 °C), with a gradual increase in large particles, which is due to the fact that at this calcination temperature interval, the amorphous silica reacts with active Al₂O₃ through Formula (11) and the amorphous silica content is low and less dissolved in the alkaline solution. These SEM observations are consistent with the leaching experiments detailed in Figure 5(c1,c2).

To obtain a more profound understanding of the microscopic morphological alterations of alkaline solution leached residues at various alkali desilication temperatures, SEM analyses were carried out, as depicted in Figure 13, the calcination temperature is 1080 °C, and the leaching conditions are 4 h, a NaOH concentration of 200 g/L, and a solid/liquid ratio of 15 g/60 mL.

In Figure 13a, the surface of the calcination clinker particles exhibits lamellar aggregates with a relatively compact architecture. Additionally, a minor fraction of fine particles (approximately 0.5–2.0 μm in size) exists as discrete entities rather than being incorporated into the aggregate structure. As the alkali desilication temperature increases from 55 °C to 75 °C in Figure 13b–d, the leached residue exhibits a granular and aggregated state, with minor changes compared to the calcination clinker, and partial aggregation of particles to a relatively large extent, about 1–3 μm. As the desilication temperature increases from 55 °C to 75 °C, the increasing leaching rate of amorphous silica leads to a gradual increase in pore structure. In Figure 13e, when the alkali desilication temperature rises to 85 °C, a large number of spherical-like zeolite P1 particles with a size of about 1.0–2.0 μm are present. In Figure 13f, when the alkali desilication temperature attains 95 °C, a substantial quantity of spherical-like zeolite P1 particles with a size ranging from approximately 2.0 to 4.0 μm becomes observable. At this temperature, the elevated crystallinity of the zeolite P1 particles results in a more perfect zeolite morphology. These SEM images are in close agreement with the XRD analysis results shown in Figure 10.

4. Discussion

In order to elucidate the desilication reaction mechanism of calcined gangue clinker, analysis of the raw materials in the coal gangue reveals the presence of two minerals, namely kaolinite and boehmite. It has been established that boehmite undergoes decomposition at a calcination temperature of 960–1120 °C, resulting in the formation of Al₂O₃ [31]. The kaolinite decomposes into meta-kaolinite at about 400–500 °C, and meta-kaolinite might decompose into mullite and amorphous silica via Formula (5) or decompose into mullite, Al₂O₃, and amorphous silica via Formula (6) at about 960–1120 °C. Theoretical assessments of the reaction equations are carried out to calculate the theoretical desilication efficiency and theoretical A/S ratios that stem from the two reaction equations. If the meta-kaolinite is decomposed according to Formula (5), the theoretical desilication efficiency achievable is 66.67%, corresponding to the A/S ratio of leached residue of 2.55. Similarly, Formula (6) yields a desilication efficiency of 75.00%, with an A/S ratio of leached residue of 3.40.

From the silica leaching efficiency shown in Figure 4, we can deduce that the reaction mechanism of kaolinite when the calcination temperature is between 960–1120 °C probably adheres to Formula (6). When compared with pure kaolin, coal gangue contains diverse proportions of boehmite. Boehmite significantly affects the A/S ratio in the alkaline leached residue products. For the purpose of understanding the influence that different amounts of boehmite in coal gangue have on the A/S ratio of the final product, an analysis of the chemical composition associated with the A/S ratio and desilication efficiency can be carried out through theoretical computations. For instance, when the proportion of boehmite relative to kaolinite is 0%, the computed alumina to silica ratio (A/S) amounts to 3.40, which is equivalent to a silica leaching efficiency of 75.00%. Additionally, if the relative content of boehmite to kaolinite is 10%, the calculated A/S ratio is 4.13, also corresponding to a silica leaching efficiency of 75.00%. In the same vein, when the boehmite content relative to kaolinite is 20%, the theoretical A/S value is 4.86, and the associated silica leaching efficiency remains at 75.00%. Theoretical evaluations indicate that as the boehmite content rises, the calculated A/S experiences a sharp increase, yet the silica leaching efficiency persistently remains at 75.00%. Significantly, the theoretical forecasts are in good agreement with the outcomes of the experimental findings.

In conclusion, the mineral transformation mechanism during coal gangue calcination and alkaline leaching is systematically elucidated through integrated experimental validation, theoretical analysis, and characterization by XRD and SEM, as illustrated in Figure 14.

From the above analysis, it can be concluded that the high-efficiency separation of alumina and silica in coal gangue within the temperature range of 960–1120 °C is due to the decomposition of meta-kaolinite. Specifically, 4 moles of meta-kaolinite break down into 1 mole of poorly crystallized mullite, along with 6 moles of amorphous silica and 1 mole of Al₂O₃. Moreover, the conversion of 2 moles of boehmite into 1 mole of active Al₂O₃ at a temperature range of 900–1100 °C also plays an important role in increasing the A/S in the leached residue. The amorphous silica is soluble in alkaline solutions, while the active Al₂O₃ and poorly crystallized mullite remain insoluble under these leaching conditions.

5. Conclusions

This paper adopts a novel method combining calcination and alkaline leaching to remove amorphous silica from high-alumina coal gangue and analyzes the reaction mechanism of desilication. The main conclusions can be summarized as follows:

(1) Increasing the calcination temperature from 900 °C to 1200 °C alters the internal phase composition of coal gangue. The distinct phase structures of alumina and silica formed under different temperatures directly influence the efficiency of silica separation. Kaolinite in high-alumina coal gangue first converts to meta-kaolinite, then decomposes into primary mullite, amorphous silica, and Al₂O₃ at 960–1120 °C; meanwhile, boehmite transforms into γ-Al₂O₃ at 940–1140 °C. The optimal desilication efficiency is achieved at the recommended calcination temperature range of 960–1120 °C, as primary mullite and Al₂O₃ remain insoluble, while amorphous silica dissolves under suitable alkaline conditions, thereby enabling efficient silica separation.

(2) The desilication temperature significantly impacts amorphous silica dissolution. Raising the temperature from 55 °C to 95 °C changes the leached residue’s phase composition. Below 70 °C, desilication is incomplete—some amorphous silica dissolves, while the rest remains solid, lowering desilication efficiency. Above 80 °C, primary mullite and Al₂O₃ react with the alkaline solution, dissolving into sodium aluminate (NaAlO₂) and sodium silicate (Na₂SiO₃) solutions, which in turn triggers the crystallization of zeolite-like phases, also reducing desilication efficiency. A suitable desilication temperature range is identified as 70–80 °C.

(3) The suitable desilication conditions involve a calcination temperature of 1080 °C, a calcination duration of 90 min, an alkaline leaching temperature of 75 °C, an alkali concentration of 200 g/L NaOH, a solid-to-liquid ratio of 1:4, and an alkaline leaching time of 4.0 h. With these conditions in place, the alumina content within the calcined coal gangue grew from 45.77% to 66.23% in the leached residue. In contrast, the silica content in the residue following leaching decreased from 51.79% to 19.35%. The silica leaching efficiency is 77.31%, the alumina leaching efficiency is 12.21%, the Na₂O content in the leached residue amounts to 1.94%, and the A/S ratio shifts from 0.88 for the coal gangue to 3.42 for the leached residue.