Chlorination Roasting Research Articles

Research on 4N8 High-Purity Quartz Purification Technology Prepared Using Vein Quartz from Pakistan

This study investigates the potential of two quartz vein ores from the Hunza District, Gilgit-Baltistan, Pakistan, as raw materials to obtain 4N8 high-purity quartz (HPQ) sand. Various quartz purification processes were examined, including ore calcination, water quenching, flotation, sand calcination, acid leaching, and chlorination roasting. Analytical techniques such as optical microscopy, Raman spectroscopy, and inductively coupled plasma spectroscopy were employed to analyze the microstructure, inclusion characteristics, and chemical composition of both the quartz raw ore and the processed quartz sand. Microscopic observation reveals that the PK-AML quartz raw ore has relatively high purity, the secondary fluid inclusions are arranged in a directional–linear manner or developed along crystal micro-cracks, and most intracrystalline regions exhibit low inclusion contents, while the PK-JTLT quartz raw ore contains a certain number of melt inclusions. The two processed quartz sand samples exhibit a smooth surface with extremely low fluid inclusion content. A comparative analysis of different purification processes shows that quartz sand calcination has a higher impurity removal rate than ore calcination. After crushing the raw ore into sand, the particles become finer with a larger specific surface area. Quartz sand calcination maximally exposes the inclusions and lattice impurity elements within the quartz, facilitating subsequent impurity removal through acid leaching. Following the processes of crushing, ultrasonic desliming, flotation, sand calcination, water quenching, acid leaching, and chlorination roasting, the SiO2 content of PK-AML processed quartz sand is 99.998 wt.%, with only a small amount of Ti and Li remaining, and a total impurity element content of 20.83 µg·g−1. This meets the standard requirements for crucible preparation in industrial applications, making this vein quartz suitable for producing high-end HPQ products. In contrast, the overall purity of PK-JTLT quartz is lower, and the high contents of impurity elements such as Li, Al, and Ti are difficult to remove via purification experiments. The SiO2 content of PK-JTLT processed quartz sand is 99.991 wt.%, which is applied to higher-quality glass products such as photovoltaic glass, electronic glass, and optical glass, thus presenting broad prospects for application.

Progress of joint process research on purification of high purity quartz sand

High-purity quartz sand is an important raw material for semiconductors, photovoltaics, aerospace and other cutting-edge fields, high-quality natural crystals can no longer meet the urgent demand for high-purity quartz in the domestic market, and it is imperative to accelerate the realisation of high-quality quartz purification technology from non-quality mineral sources. This paper summarises the joint process flow in the unused scenario, and outlines the main methods for the purification of quartz sand in China at present, including pretreatment, ultrasound-assisted leaching, chlorination roasting, etc., from the point of view of the joint process, and puts forward the problems that exist in these methods, so as to provide certain references for the research on the purification process of high-purity quartz sand in China.

Chloridizing roasting studies of spent NdFeB magnets for recovery of rare earth values

The increased demand for rare earth elements in advanced technological applications and supply shortages call for metal recovery from secondary sources. Permanent magnet (Nd2Fe14B or NdFeB) may serve as a potential secondary source due to its high rare earth (Nd+Pr+Dy: ∼30 %) content and its vast application. The present study utilizes a chloridizing roasting (CaCl2.2 H2O) pre-treatment process followed by water leaching, acid leaching (0.5 M HCl, S/L =1/10 g/ml, 90 °C, 3 h), oxalic acid precipitation and calcination (850 °C, 2 h) to obtain mixed rare earth oxides. The process was optimized based on temperature (400–700 °C), dosage (CaCl2.2H2O: NdFeB=0.5:1–2.5:1), and time (30–120 min) on the rare earth dissolution. The theoretical activation energy for the chloridizing roasting process is estimated as 22.3 (OFW) and 16.7 kJ/mol (KAS), while the experimental activation energy for Nd and Dy dissolution was determined to ∼29.3 and ∼17.7 kJ/mol, respectively depicting product layer diffusion-controlled kinetics. Higher dosages of CaCl2.2H2O (1.5:1 and 2:1) favored NdOCl formation, thereby, higher dissolution; however, further higher dosage (2.5:1) leads to reduced Nd dissolution due to higher CaO formation and acid consumption by Ca during leaching. Incomplete oxidation at lower temperatures (400 °C) and iron dissolution impair the Nd dissolution and selectivity. Excessive oxidation at >700 °C favors the formation of NdFeO3, decreasing Nd dissolution. The maximum dissolution of Nd was ∼89 %, while for Dy, it was ∼88 % at optimum conditions of 600 °C, 90 min, 2:1. Water leaching post-roasting leads to ∼87 % Ca removal and the precipitation efficiency of rare earth oxalates was 99 %. The overall extraction for rare earth elements was ∼89 %, and 1 kg of NdFeB powder can yield ∼285 g of rare earth oxides (∼239 g Nd2O3, ∼14 g Dy2O3) with 96 % purity. Further, this study demonstrates that using CaCl2.2 H2O as a solid chlorinating agent in chlorination roasting enhances recovery rates of mixed rare earth oxides while providing a safer and more environment-friendly alternative for industrial applications.

Novel Strategy for Efficient Recovery of CuO-Based Multiple-Metals from Copper Smelter Dust toward CO2 Electrocatalytic Reduction.

Copper smelter dust, a typical hazardous waste that is abundant in valuable heavy metals, holds the potential to be regarded as a promising resource. This study introduces a new approach that integrates chlorination roasting and cascade condensation to efficiently recover heavy metals from copper smelter dust. The findings demonstrate the successful separation of heavy metals (Cu, Pb, and Zn) as chlorides at nearly 100% efficiency while also effectively converting trivalent arsenic (As(III)) into pentavalent arsenic (As(V)) and immobilizing it in the roasting residues, thereby reducing environmental risk. Through the utilization of thermogravimetric mass spectrum analysis and thermodynamic equilibrium calculations, the chlorination process for heavy metals was investigated, revealing both direct and indirect chlorination processes. Additionally, the study resulted in the development of a CuO-based multiple-metals electrocatalyst from the oxidized roasting-recovered heavy metal chlorides, exhibiting significantly enhanced catalytic activity and faradaic efficiency for the electroreduction of CO2 into CO and CH4 compared to pure CuO electrocatalyst under similar electrocatalytic conditions. Overall, this work presents a sustainable and scalable method and new insights for addressing environmental risks while repurposing copper smelter dust.

Preparation of Oxidized Pellets from Sulfuric Acid Residue Containing Zinc and Lead by Chlorination Roasting and Its Mechanism of Dezincing and Lead Removal

The utilization of sulfur acid residue is an urgent problem confronting sulfuric acid production enterprises, especially the application of sulfur acid residue (SAR) containing lead and zinc. A method combining chlorination roasting using CaCl2 with pelletizing for processing SAR containing lead and zinc was used in this study, and the effect of calcium chloride on pelletizing performance was studied; in addition, the removal behavior of lead and zinc was also studied by using polarized light microscopy (Zeiss double Axioskop 40A), X ray diffraction, SEM and EDS. The results showed that CaCl2 migrated to the surface of the pellets during drying, and this phenomenon resulted in a lower removal rate of lead and zinc inside the pellets than outside the pellets during the preheating phase. When the roasting temperature was 1220 °C, with an increase in the basicity of pellets, the silicate minerals in the pellets gradually decomposed, the hematite particles were gradually refined, and more lead or zinc minerals were exposed, which further increased the removal rate of lead and zinc in the pellets. Finally, the SAR pellets with Pb and Zn removal rates up to 91.33 and 97.88%, and a compressive strength of 2789 N, could be obtained, which is very beneficial to the sustainable development of sulfuric acid mills.

Preparation of High-Purity Quartz Sand by Vein Quartz Purification and Characteristics: A Case Study of Pakistan Vein Quartz

This study focuses on the purification and evaluation of the high-purity quartz (HPQ) potential of vein quartz ore from Pakistan. Vein quartz is grayish-white and translucent, with its mineral composition mainly comprising quartz crystal. Processed quartz sand is obtained from quartz raw ore through purifying technologies, including crushing, ultrasonic desliming, flotation, high-temperature calcination, water quenching, hot pressure acid leaching, and chlorination roasting. The microscopic characteristics show that the vein quartz raw ore has a medium-coarse granular metacrystalline structure, high quartz content, with only a small quantity of fine-grained K-feldspar. The inclusions primarily consist of large-sized primary inclusions and secondary fluid inclusions developed along the micro-fractures, and the content of inclusions in most areas of the crystal is very low or even nonexistent. The quartz ore with such inclusion characteristics is considered a relatively good raw material for quartz. Component analysis shows that the main impurity elements in the quartz ore are Al, K, Ca, Na, Ti, Fe, and Li, with a total impurity element content of 128.86 µg·g−1. After purification, only lattice impurity elements Al, Ti, and Li remain in the processed quartz sand, resulting in a total impurity element content of 24.23 µg·g−1, an impurity removal rate of 81.20%, and the purity of SiO2 reaching 99.998 wt.%. It is suggested that when the quartz raw ore contains high content of lattice impurity elements, such as Al, Li, and Ti, it is difficult to remove them by the current purification method. In industrial production, considering the economic cost, if quartz sand still contains high content of lattice impurity elements Al, Ti, and Li after flotation, it cannot be used as a raw material for high-end HPQ.

Behavior of gold and non-ferrous metals during low-temperature chlorination roasting of e-waste

Behavior of gold and non-ferrous metals during low-temperature chlorination roasting of e-waste

Thermochemically driven crystal phase transfer via mechanical activation-assisted chlorination roasting toward the selective extraction of lithium from spodumene

With the rapid advancement and widespread application of lithium-ion batteries, the demand for lithium experiencing a precipitous surge. Hence, there is an increased emphasis on lithium extraction from spodumene deposits. However, the sustainable expansion of the lithium industry encounters challenges. Conventional processes suffer from drawbacks such as elevated transformation temperatures, high acid consumption, and challenges in leachate purification. In response to these challenges, this work proposes a selective and efficient lithium extraction process from α-spodumene through the combined approach of mechanical activation-assisted chlorination roasting and water leaching. Utilizing density functional theory (DFT) calculations to tailor chloride salts for the thermal treatment of spodumene, the combination of thermal behavior characterization and thermochemical calculations indicates the favorable propensity of CaCl2 and its hydrates for preferential reactions. Through the amalgamation of mechanical activation and chlorination roasting, the outcomes reveal that mechanical activation could reduce the phase transition temperature. Following a 3-hour activation period, a complete chlorination reaction is achievable at 900℃, yielding a lithium extraction efficiency of 91.34%. This endeavor holds the potential to furnish novel insights into the mechanism of selectively extracting chlorination roasting, thereby offering promising avenues for developing environmentally sound and selectively efficient technologies for lithium extraction from spodumene.

Effect of Chlorinating Agents on Lithium Recovery and Product Purity from Lithium-ion Battery (LIB) Recycling Slags via Chlorination Roasting Process

ABSTRACT With the rapid growth of the electric vehicle (EV) and energy storage system (ESS) markets, the demand for lithium-ion batteries (LIBs) has skyrocketed, highlighting the importance of securing a sustainable supply of lithium. This study explores chlorination roasting as a method for recovering lithium from Li-containing slag, which is a promising alternative lithium source obtained from pyrometallurgical LIB recycling process. We investigate the efficiency of chlorination roasting in producing technical grade (TG) lithium chloride (LiCl) from the slag and quantitatively analyze the impurity behavior during the process. Different chlorinating agents, including CaCl2, NaCl, and PVC, are evaluated for their effectiveness in lithium recovery and purity. The results indicate that CaCl2 is the most effective in achieving high lithium recovery and LiCl purity, at approximately 90%. This offers a promising solution for sustainable lithium production. This research would highlight the importance of developing efficient recycling methods to meet the growing demand for lithium in EV and ESS applications.

Chlorination mechanism of typical metallic radionuclides by HCl on irradiated graphite surface from first principles

Radioactive activity of metallic radionuclides accounts for a considerable proportion in irradiated graphite from nuclear reactors. Chlorination roasting is a promising method to remove metallic radionuclides in decommissioned irradiated graphite. The decontamination effect and chlorination mechanism of 90Sr and 60Co by HCl over an irradiated graphite surface were estimated by density functional theory. 90Sr and 60Co could be chlorinated via X(signifying metallic radionuclide)→XCl→XCl2 reaction pathway. The rate-determining step appeared where metallic radionuclides desorbed from the irradiated graphite surface. The chlorination energy barriers of 90Sr and 60Co were 236 and 575 kJ/mol, respectively, implying that chlorination roasting method for decontaminate of 90Sr and 60Co in the irradiated graphite was feasible. This work is beneficial to the application of chlorination roasting for removal of metallic radionuclides in irradiated graphite.

Study on mechanism and kinetics of iron removal by chlorination roasting of coal gangue

Utilizing the abundant kaolin resources within coal gangue is economically beneficial, although the iron-containing phase in the minerals hampers the value of its utilization. In the present study, CaCl2 served as the chlorination agent for iron removal from coal gangue through chlorination roasting. The study revealed that when the roasting temperature is below 800 °C, CaCl2 decomposes into HCl and reacts with the iron content to form chloride, which then volatilizes. Comparatively, when the roasting temperature exceeds 800 °C, CaCl2 will decompose to produce HCl and Cl2 to accelerate the volatilization of iron. Through non-isothermal kinetic analysis, the apparent activation energy for iron removal by chlorination roasting of coal gangue was found to be 82.42 KJ/mol, and the pre-exponential factor was 6329.3 s−1. Correspondingly, isothermal kinetic analysis revealed the apparent activation energy of 84.2 KJ/mol and a pre-exponential factor of 6310.7 s−1. The optimal reaction kinetic mechanism function for iron removal through chlorination roasting of coal gangue was identified as the F3 chemical reaction order model via above kinetic analysis. The integral took the form g(α) = (1-α))2–1.

An efficient approach to utilize copper smelting slag: Separating nonferrous metals and reducing iron oxide at high temperature

Over 50 million tons of copper slag are produced worldwide annually. Stacking is currently the primary method used to treat copper slag, resulting in resource wastage and environmental issues. Using slag as a raw material in the steel industry is areasonablesolution. However, the presence of nonferrous metals degrades steel performance and corrodes smelting equipment, and the remaining slag poses environmental and sustainable challenges. Thus, this study focused on removal of nonferrous metals from copper slag and subsequent reduction of iron oxide. The experimental results showed that increasing the percentage of the chlorinating agent (calcium chloride), temperature, and duration, the removal percentage of copper initially increased linearly, then plateaued. The acidity coefficient had a marginal effect on copper removal percentage. The optimum chlorination roasting conditions were 13 % calcium chloride, 1373 K, and 0.5 h, resulting in removal percentages of 90.3 %, 81.9 %, and 82.7 % for copper, zinc, and lead, respectively. The appropriate oxygen partial pressure for chlorination roasting was 10-5–0.7 atm. The reduction percentage of iron oxide was approximately 82 % under roasting conditions of 1.5 h at 1373 K. Based on these results, a novel scheme for copper slag utilization is proposed that involves chlorination roasting for nonferrous metal removal, reduction roasting for iron oxide reduction, and melting to obtain molten iron and separate molten slag. The iron and slag are used for steel manufacturing and rock wool preparation, respectively. This scheme provides apromisingway to efficiently use copper slag, which will reduce primary resource consumption and pollution.

Investigation on decontamination kinetics of metallic radionuclides in irradiated graphite by chlorination roasting

Irradiated graphite used as moderator and reflector material in reactor core in nuclear industry is an important radioactive waste after reactor decommissioning. Treatment of irradiated graphite waste has been a challenge for IAEA member states. Chlorination roasting was proposed in order to remove the metallic radionuclides in irradiated graphite waste. The promising advantages of the proposed method include that the category of graphite waste can be reduced with near-zero carbon emission, and it can lead to potentially the recycling of irradiated graphite and metallic radionuclides. The decontamination kinetics of metallic radionuclides was evaluated based on a lab-scale tube furnace combined with density functional theory (DFT). 56Fe ions were implanted into nuclear graphite to simulate metallic radionuclides derived by neutron radiation. Decontamination efficiencies of Fe at 900, 1000, 1100, and 1200 °C for the retention time of 210 min were 42.6%, 81.6%, 94.4%, and 96.7%, respectively. Fe could volatilize from graphite in the form of complex metal chloride with retaining a large amount of graphite inert matrix. The decontamination process was described by the second order reaction model. Through DFT evaluation, Fe was chlorinated via Fe → FeCl → FeCl2 reaction pathway. This decontamination process is efficient, cost-effective, and reliable, which provide a valuable contribution to managing irradiated graphite.

Recovery of LiCl and Co3O4 from the cathode material contained in spent lithium-ion batteries using chlorination roasting with MgCl2·6H2O

Highly demanded lithium and cobalt are considered critical metals because their reserves in the earth’s crust are relatively low in abundance, and they are found in a few countries. Thus, the extraction of these metals from spent lithium-ion batteries (LIBs) has gained much attention since they are present in relatively high concentrations in the cathode of the battery. Chlorination roasting is an attractive technique of metal extraction because of its salient features: short reaction times and moderate reaction temperatures, selectivity to the extraction of certain metals and in many cases implies the use of low-cost chlorinating reactants. However, there are few studies concerning the recycling of metals through chlorination roasting. This research paper evaluates the chlorination roasting using MgCl2·6H2O as chlorinating reactant followed by water-leaching to recycle lithium and cobalt from the cathode material of spent LIBs. First, a non-isothermal chlorination test between 20 and 900 °C was performed to evaluate how temperature affects the chlorination of the metals and the presence of impurities. Then, isothermal chlorination tests at 420, 440 and 470 °C for times between 60 and 350 min were performed to evaluate the effect of time on the purity of the product. It was found that the chlorination treatment is selective to lithium. HCl(g) generated during the decomposition of MgCl2·6H2O reacts with lithium to produce LiCl. Cobalt is extracted as Co3O4, as it is another product of the chlorination reaction. The product of LiCl impurified with Mg(OH)Cl, formed during the extraction process, was purified by a heat treatment in air at 570 °C for 60 min followed by water-leaching. This treatment also generates MgO as a by-product. The optimum conditions of the extractive process were the followings: 440 °C and 210 min. The extraction yields of lithium and cobalt were 83.29% and 89.66%, respectively. The recycling process proposed here proved to be efficient. Moreover, the magnesium chloride generated as waste from the lithium extraction process from brines could be applied in the process from a circular economy point of view.

The Recovery of Gold in Low-Grade Cyanide Tailings by Chlorination Roasting

The resource utilization of cyanide tailings has significant environmental and economic benefits. The efficient recovery of gold from low-grade cyanide tailings containing 1.71 g/t Au was performed by a chlorination roasting–flotation process. The effects of roasting temperature, calcium chloride, internal coke, external coke, copper sulfide concentrate, and kaolin on the recovery rate of concentrate, gold grade, and sorting efficiency were investigated. The optimized process parameters were as follows: 16 wt% calcium chloride dihydrate, 6 wt% internal carbon, 1 wt% external carbon, 9 wt% copper sulfide concentrate, 2 wt% kaolin, and roasting temperature of 730 °C. The sorting rate, gold grade, and recovery rate of gold concentrate can reach 88.48%, 33.46 g/t, and 76.7%, respectively, and the gold grade of tailings was as low as 0.17 g/t. In the matte phase, gold can be enriched in the form of gold grains. Therefore, through chlorination roasting, the trapped gold can be released and deposited on the surface of the matte phase due to the transformation from hematite to magnetite. The gold-deposited metal sulfide can be effectively recycled through flotation. These results have potential guiding significance for the efficient recovery of gold from cyanide tailings.

Potassium chloride recovery from mechanically activated microcline through the chlorination roasting and leaching route

In this study, effects of mechanical activation in the chlorination roasting and water leaching route known as CaCl2 process and developed for the production of potassium chloride (KCl) from potassium feldspar ores were studied. A microcline containing K-feldspar ore with 10.89% K2O was first intensively dry milled by a planetary ball mill and mixed with calcium chloride (CaCl2) and then roasted at temperatures up to 1000 °C to obtain KCl that will be finally dissolved by the water leaching. Potassium recovery by water leaching increased rapidly up to 800 °C. At higher temperatures, the recovery decreased fast due to the evaporation of KCl. According to the K recovery values per unit energy consumed, the optimum roasting temperature was determined as 750 °C and the milling time was 15 min. It was concluded that intensive milling causes mechanical activation of the microcline to reduce the chlorination roasting temperature, which triggers a rise in the K recovery by the water leaching.

Preparation of a High-Silicon ZSM-5 Molecular Sieve Using Only Coal Gangue as the Silicon and Aluminum Sources.

The traditional preparation of ZSM-5 molecular sieves relies on chemical reagents to provide silicon and aluminum sources, which are limited as raw materials and cannot be commonly used in production practice. Using coal gangue as the raw material and using medium-temperature chlorination roasting and the pressure acid leaching process to control the silicon-aluminum ratio [n(Si/Al)] of coal gangue, a ZSM-5 molecular sieve was prepared using the alkali melting hydrothermal method. The pressure acid leaching process solved the limitation that kaolinite and mica cannot simultaneously be activated. Under optimal conditions, the n(Si/Al) of the coal gangue increased from 6.23 to 26.14 and complied with the requirements for the synthesis n(Si/Al) of a ZSM-5 molecular sieve. The effect of n(Si/Al) on the preparation of the ZSM-5 molecular sieve was studied. Finally, spherical granular ZSM-5 molecular sieve material with a microporous specific surface area of 169.6329 m2/g, an average pore diameter of 0.6285 nm, and a pore volume of 0.0988 cm3/g was prepared. Providing ideas for the high-value utilization of coal gangue, it is significant for solving the problem of coal gangue solid waste, as well as the problem of ZSM-5 molecular sieve feed stock.

Elimination of iron in molybdenum concentrate via vacuum chlorination

Due to the great affinity and the similarity of the physical performance of iron and molybdenum disulfide in the concentrate, it is difficult to affect the removal and separation of iron during the procession of molybdenum concentrate. The application of molybdenum disulfide two-dimensional layered materials is seriously restricted by the relatively high content of iron. Considering the direct purification of MoS2, vacuum chlorination roasting for decreasing iron of molybdenum concentrate is proposed. Thermodynamics of iron removal by vacuum chlorination roasting is analyzed. FeS can be chloridized into FeCl2, which volatilizes easily by solid chlorination agent in vacuum. The experimental results showed that the iron content decreased to 0.43% at 1223 K and P ∼1 Pa for 60 min, and the Fe content was reduced to 0.082% after acid washing, with a removal rate of 94.2%. which offers a novel way to replace the conventional method for elimination of iron sulfide in molybdenum ore. Possible separation mechanism of iron sulfide is also discussed in this work.

Mechanism and kinetics for chlorination roasting of copper smelting slag

An efficient process based on chlorination roasting with CaCl2 was proposed to recover zinc from copper smelting slag. The reaction mechanism and the kinetics of the chlorination process were investigated by thermodynamic calculation, thermogravimetry and differential scanning calorimetry (TG−DSC) analysis and X-ray diffraction (XRD) methods. The results demonstrated that the temperatures of oxidative decomposition of CaCl2 and chlorination of all Zn-containing phases were above 774.3 and 825 °C, respectively. The chlorination roasting process was divided into four stages: precipitation of adsorbed water, extraction of crystallized water, oxidation of Fe-containing phases, and chlorination volatilization of zinc. The average apparent activation energies of the iron oxidation and zinc chlorination were 101.70 and 84.4 kJ/mol, respectively. The most probable mechanism function for the iron oxidation process was the Avrami−Erofeev model (n=2). Zinc chlorination followed the shrinking unreacted core model, and the chemical reaction was the rate-controlling step.

Behavior of the components of carbon-containing ash from the combustion of power coals under the conditions of chlorination roasting

On the basis of thermodynamic calculations of interaction reactions between mullite, oxides of titanium, iron, and Cl2 (gas) in the presence of carbon, the values of the Gibbs free energy are determined in the current work. It has been established that the probability of reactions occurring in the temperature range 1073–1473 K is very high. The most negative value ΔG°Т = –1029.9 kJ/mol is typical for the titanium oxide reduction reaction at the roasting tem- perature (1100 °С). The value of the Gibbs free energy of the destruction reaction of mullite (Al2SiO5), which is important for practice, is ΔG°Т = –269.2 kJ/mol. This makes it possible to predict the complete decomposition of mullite from ash to easily soluble aluminum chloride under the conditions of chlorination roasting. The study of the behavior of carbon under the conditions of chlorination roasting of ash with CaCl2 in an oxidizing atmosphere is also important. The results of the differential thermal analysis of the non-magnetic fraction of ash, previously separated from the TPP ash by magnetic separation, showed clearly defined manifestations on the DTA, DTG and TG curves, caused by the presence of various types of reactions in the system. Low-temperature (20–115 °C) sample dehydration and ash dehydration were determined in the temperature range of 115–375 °C. In the range of further heating of the sample (375–685 °C), an exothermic projection was found on the DTA curve, which is characteristic of the oxidation of thermally active carbon present in the powder. The amount of carbon removed from the ash established as 6.75% by weight of the sample. The research was carried out within the framework of grant funding from the Committee of Science of the Ministry of Education and Science of the Republic of Kazakhstan for 2021–2023 in the priority area “Geology, mining and processing of mineral and hydrocarbon raw materials, new materials, technologies, safe products and structures” of project No. АР09259637 “Development of a highly efficient wastefree technologies for the utilization of ash from coal combustion with the production of marketable products.