In order to study the mechanical mechanism of the instability of the reverse fault in the working face under the condition of thick and hard roof, the spatial relationship model between the overburden movement area and the fault plane was established through the relationship among the size of the coal pillar at the fault boundary, the fault dip angle and the angle of rock strata movement. The equation for the movement line of overlying strata was established based on the movement angle of each rock layer in the mining area. The distance function formula between the movement line of different layers of overlying strata and the fault plane was obtained by combining the fault dip angle, and the boundary transition area between the fault plane and the rock layer movement line was accurately delineated. A limit equilibrium equation for boundary rock blocks was established based on the tensile strength of rock layers, and a mechanical instability criterion for boundary rock blocks was given when the fault dip angle was less than the rock layer movement angle. Based on on-site cases, the activation rock zone of reverse faults with a dip angle less than the rock movement angle was determined using dual criteria, and the activation law of fault planes and the spatial distribution of separation layers were determined. The deformation and failure mechanism of sur-rounding rock under the boundary condition of reverse fault is theoretically analyzed. The overall migration law of rock strata in the fault area is simulated by experiments. The mechanical and displacement variation modes of surrounding rock in the reverse fault area under the condition of thick and hard roof are comprehensively calculated and explored. The results show that the reverse fault is activated before the breaking of the thick and hard roof. When the fault is less than the dip angle of the rock stratum, the separation space of the fault surface is mainly distributed in the bending subsidence zone of the thick and hard roof area. The vertical fracture space of the rock stratum which is separated from the fault plane is significantly reduced. the upward transfer efficiency of the rock layer that generates the separation space with the fault surface to the goaf space is significantly improved. The stability of fault surrounding rock can be maintained by optimizing the size of fault protection coal pillar. The double criterion model gives the size value basis of the fault boundary protection coal pillar from the theoretical level, and the optimal solution of the field protection coal pillar width is provided.

The key to safe and efficient mining of steeply dipping coal seam is the multi-dimensional coupling control of surrounding rock and equipment. Quantitative characterization of the spatial structure fracture characteristics of the overlying rock, the development of mining fractures, and the distribution characteristics of voids in the mining area are the prerequisites for achieving precise rock layer control in such coal seams. Based on the engineering background of steeply dipping longwall and large mining height working face, the research methods of numerical calculation, field measurement and theoretical analysis are combined. On the basis of determining the spatial movement and deformation law of overburden rock, the three-dimensional curvature, rock porosity, and fractal geometry theory were introduced to quantitatively describe the migration, stacking, and hinge characteristics of fractured rock blocks in different areas of the mining site, as well as the distribution law of rock fractures, Realized the dynamic and precise quantitative characterization of the three-dimensional structural composition and spatial occupancy of mining induced fracture fields during the mining process of steeply dipping and high mining height working faces. Research has shown that the fractured roof with steeply dipping and high mining height undergoes multi-dimensional coupled non equilibrium movement in the mining space. The middle and upper roof mainly moves vertically, while the lower roof evolves into a inclined downward movement. The form of roof failure presents three-dimensional regional differentiation along the dip, strike, and vertical directions. The upper broken rock blocks are staggered and migrated across layers to form a masonry structure. Behind the goaf, there is a “non-uniform gangue – curved cantilever beam – broken rock block” bearing structure that forms a non-uniform empty roof area. The bearing structures are constrained by each other and undergo dynamic evolution with different dip positions and overlying rock layers. The support resistance is distributed in fluctuating zones, with the middle resistance being greater than the upper and lower parts, and the strike is not synchronized with the pressure. The curvature distribution shows a “horizontal and vertical O–X” shape, and the “O” shaped internal collapsed rock blocks sink non parallel and synchronously. The inclined and anti inclined stacking structures coexist, and the positive and negative curvature sizes vary, forming weak articulated structures between the rock blocks. As the overburden layer increases, the detachment failure area shifts towards the middle and lower parts and the middle part of the goaf, and the porosity and fractal dimension of the rock mass first increase and then decrease. The research results have revealed the three-dimensional fracture movement law of the overlying rock within the goaf, enriched the control theory of the steeply dipping coal seam strata, and provided a new method for precise quantitative analysis of the movement law of the roof strata.

The study was conducted on modification of grouting of broken roof under stratified mining in a steeply dipping coal seam. Considering the geological conditions of stratified mining in the merged coal seams 3 and 1 at the Pansidong Mine, five types of particle size gradation of grouting reinforcement body models were established, a single-axis compression test based on CT in situ scanning was carried out, and the effects of particle-sized paired reinforcement body strength characteristics, crack evolution, acoustic emission signals and micromorphological characteristics were studied. The study showed that: ① With the increase of particle size grading, defects such as pores and cracks in the reinforcement body are affected by the dislocation of the rock and plasma matrix. Large-scale shear belts have been formed, leading to a gradual decrease in the single-axis pressure resistance of the reinforcement body. The 6～8 mm and 8～10 mm particle size grades of reinforcement body are affected by large particle size rocks, and early stress fluctuations are evident. The form of destruction shows a transition from cross-strait destruction, a combination of slash-slash destruction and a slash destruct. ② With the increase of particle size gradation, the hydrated products (AFt and C–H–S) do not form interlaced network structure to fill the pores between the blocks, which leads to the decrease of the overall strength of the reinforcement body. During the loading process, the end of the rock generates concentrated stress, the cut band increases, the sound emission activity is earlier and more frequent, the high-value vibration count increases and the cumulative vibration count grows faster. ③ The CT scan shows that the main fracture of the reinforcement body occurs mainly at the cementation surface between the rock mass and the grout matrix, and rarely extends through the rock. With the increase of particle size gradation, more small cracks in the reinforcement body expand into large opening cracks, and the number of cracks decreases. After the destruction of 0～2 mm, 4～6 mm and 8～10 mm particle sizes of reinforcement body, the proportion of the crack volume V ≤ 0.01 mm³ decreased by 9.5%, 1.6%, 4.0%, and the proportion of cracks V > 10 mm³ increased by 0.1%, 3.5%, 4.9%. The graying value of the crack increases, and the proportion of the crack volume increases to form a more complex crack network. Based on the test results, the drilling and grouting construction and netting scheme of broken roof are proposed, which provide a theoretical basis for the stable control of broken roof under stratified mining in a steeply dipping coal seam.

The overlying strata fracture caused by shallow and close coal seam mining is developed, and air leakage condition is complicated, which leads to prominent hidden danger of residual coal spontaneous combustion and seriously restricts the safe and efficient production of Lijiahao coal mine. Therefore, similar experiment and numerical simulation were used to delimit mining-induced fracture zone in overlying strata of shallow buried compound goaf, fractal geometry theory was utilized for quantifying the fracture in zone of overlying strata. Based on the key layer theory, the cause of the overlying strata fracture zone was revealed and targeted control technology plan was proposed. The results showed that, the overlying strata fracture circularly opened and closed during 31114 working face mining, while the surface periodically produced dip dynamic crack behind the working face. Ultimately, the fracture distribution in overlying strata of shallow buried compound goaf presented “M shaped”, which could be divided into irregular fracture zone Ⅰ in caving zone, oblique breaking fracture zone Ⅱ, Ⅲ and oblique separation fracture zone Ⅳ in fracture zone and the edge crack was formed inside the goaf boundary. Among them, the void of fracture zone Ⅰ intricately developed and disorderedly distributed, so the fractal dimension of the fracture network was the largest, 1.569. Then followed by fracture zone Ⅱ, Ⅲ and Ⅳ, 1.531, 1.396 and 1.438 respectively. The mechanical model and instability criterion about “voussoir beam” structure of broken key rock block in overlying strata of shallow buried compound goaf were built to explain that the rotation deformation and instability of the “voussoir beam” structure of key layer in main roof led to the oblique breaking fracture opening in fracture zone Ⅱ and Ⅲ, while the combination bearing effect of the “voussoir beam” structure of key layers in main roof and interlayer resulted in oblique separation fracture development in fracture zone Ⅳ. The surface-underground collaboration control technology of overlying strata oblique fracture and surface crack in shallow buried compound goaf was put forward. That was the advancing speed of the working face was increased to 14 m/d at the stage of entering and leaving the overlying goaf and final mining, combined with backfilling wind-deposited sand and covering loess to deal with surface crack of fracture zone Ⅱ, Ⅲ and Ⅳ. SF₆ tracer technique test proved that the surface air leakage was small and the control effect of overlying strata mining-induced fracture in shallow buried compound goaf was good, which provided an important theoretical basis for the prevention and control of residual coal spontaneous combustion.

Addressing significant deformation in the surrounding rock of deep soft rock road-ways subjected to repetitive repairs, this study focuses on the 21st transportation downhill of Quandian coal mine as the engineering context. We propose the variable-diameter partition roadway pressure relief technology. This approach utilizes theoretical analysis, numerical simulation, and practical engineering methods. We established a theoretical model for pressure relief parameters, derived essential parameters for pressure relief, and outlined the distribution pattern of elastic strain energy density. Additionally, we uncovered the principles governing energy dissipation in borehole unloading, ultimately determining the optimal value for key pressure relief parameters. The study findings indicate the following: ① The variable-diameter zoned pressure relief technology enlarges the energy reduction zone near the roadway while diminishing the energy elevation zone in the roof and floor. It redirects concentrated energy from the shoulder and floor corners deeper into the area, achieving precise pressure relief control within the roadway. ② Increasing the length of shallow small-diameter boreholes (L₁), the length of deep large-diameter boreholes (L₂), and the radius of deep large-diameter boreholes (r) corresponds to increasing pre-peak, post-peak, and post-peak values in the elastic strain energy density curves, demonstrating an overall positive correlation. ③ As the spacing (D) of deep large-diameter boreholes increases, the peak value of elastic strain energy density between holes shows the characteristics of firstly increasing, then decreasing and then stabilizing, and the effect of unloading joint pressure between boreholes is negatively correlated with it, and based on the magnitude of the increase in the peak value of elastic strain energy, the key parameters affecting the effect of unloading are weighted as follows in the order of: L₁, L₂, r and D. ④ Based on the theoretical analysis and numerical simulation to determine the key parameters of the 21st transportation downhill roadway pressure relief technology reasonable value, roadway implementation of variable-diameter zoned pressure relief technology, the two ribs and the roof and floor deformation were reduced by 58.7%, 23.7%, and 27.4%, respectively, the stability of the roadway has been effectively controlled, confirming the effectiveness of the variable-diameter zoned pressure relief technology.

During the process of mining of the steep and ultra-close multiple coal seam groups, the roadway is affected by multiple mining, the regional stress environment is complex, the deformation and failure of the roadway is severe and the non-uniform characteristics are significant, and the support and maintenance are difficult. For such problems, taking the repeated mining roadway of the and ultra-close multiple steep coal seam 31233 transportation lane of Daichiba as the engineering background, the comprehensive research methods such as theoretical analysis, numerical simulation and field experiment are adopted to find out the evolution characteristics of surrounding rock stress environment at different locations of the roadway and the dis-similatory failure law of roadway surrounding rock under its influence, the special-shaped section which is beneficial to the convergence of plastic failure distribution of surrounding rock is designed, and the design method of non-uniform bolt ( cable ) support based on the alienation distribution law of plastic zone of roadway surrounding rock is proposed. The results show that under the influence of repeated mining of the steep and ultra-close multiple coal seam, the bidirectional principal stress around the roadway has a large difference, and the direction of the principal stress is deflected accordingly, resulting in the butterfly-shaped alienation characteristics of the surrounding rock of the roadway, such as non-uniformity and deflection of the maximum failure depth. The stagger position and cross-section shape of the roadway determine the failure form and scope of the roadway, and the stagger distancet directly affects the dissimilation distribution and scope of the plastic zone of the surrounding rock, and the existence of hard rock causes local annihilation of the plastic zone. When the position of the roadway is staggered to more than 17 m, the stress environment around the roadway tends to be stable, and there is no obvious change in the plastic size and shape of its surrounding rock. The internal dislocation position is greater than the distance, which is most favorable for the stability of the surrounding rock of the roadway. The unilateral span size of the roadway section determines the depth of plastic failure on its side. When the roadway section is a polygonal special-shaped section, the convergence degree of plastic distribution of surrounding rock is significant. In the design of roadway section, the longest side of roadway should be located on the side of hard rock stratum and try to avoid the direction of maximum failure depth, so as to make full use of the annihilation characteristics of plastic zone of hard rock stratum and avoid the excessive expansion of butterfly leaves in plastic zone of roadway surrounding rock. Therefore, the geometric characteristics of the section of the special-shaped roadway with the condition of steep coal seam can be fully utilized. Based on the dissimilation distribution law of the plastic zone of the surrounding rock of the roadway and the distribution characteristics of the reasonable anchorage layer, the non-uniform bolt (anchor cable ) support design can be carried out, which can effectively guarantee the support force of the bolt (anchor cable ) and ensure the overall stability of the surrounding rock of the roadway. At the same time, the industrial test was carried out in 31233 headentry. Before the influence of severe mining, the monitoring station of deep displacement of roadway roof and anchor cable support force was set up. During the monitoring period, the total deformation of roof deformation was controlled within 100 mm, the anchor cable support force was stable, and the overall control effect of surrounding rock was better.

China's coal resources with a buried depth of more than 1000 m account for more than 50% of the total reserves, which will be the main part of coal resources supply in the future. It is of great significance to carry out the scientific productivity of rock burst mines and scientifically improve the productivity to ensure the safe and efficient mining of deep coal resources in China. Aiming at the problem of productivity determination and productivity improvement of rock burst mine, the range of mine field containing rock burst coal seam is scientifically divided by theoretical analysis and field monitoring, and the disturbance response characteristics of mining face under different mining methods, mining sequences and mining layouts in rock burst mine are analyzed, the determination method of scientific productivity of rock burst mine is established, and the top-level design and practice method of scientific productivity improvement of rock burst mine are put forward. The research results show that it is necessary to identify the impact tendency and determine the impact risk of the rock burst mine, and to realize the scientific planning of the mine field range with rock burst coal seam; The mining method, mining sequence and layout, and the response characteristics of mining parameters to the disturbance of coal face are studied, and the systematic method of scientific design of coal face in rock burst mine is obtained, which creates conditions for the productivity improvement of rock burst mine. This paper puts forward the technical path to determine the productivity of rock burst mine, and based on mining dynamics, constructs a method to determine the advancing speed of coal face in rock burst mine by stages, regions, periods, unloading pressure and historical prevention and control level. The top-level design of productivity improvement in rock burst mine is constructed, and it is clearly pointed out that the key to control productivity is the impact risk brought by productivity improvement, so as to achieve the balance between mining intensity and unloading pressure and reduce the impact risk, thus effectively improving productivity. The integrated technical path from scientific planning of mine field scope and scientific design of coal face to scientific determination and improvement of productivity in rock burst mine is constructed, which provides scientific guidance for scientific determination and improvement of productivity in rock burst mine.

In the 3D reconstruction of coal-rock combinations fractures, in response to the problem that traditional threshold segmentation methods cannot accurately determine the threshold size between coal and rock, resulting in poor fracture segmentation performance, a new VRA-UNet coal-rock combinations fracture identification model based on deep learning theory is proposed, providing an optimized solution for accurate identification of coal-rock combinations fractures. Firstly, the VGG16 module is used as the backbone feature extraction network to enhance the model’s generalization ability and prevent the initialization of model parameters from being too random. Secondly, to address the complex fracture topology and strong non-uniformity of coal-rock combinations, an attention module (ResCBAM) with spatial and channel dimensions is introduced into the up-sampling part to enhance the model's feature extraction ability and alleviate the problem of gradient disappearance. Finally, an asymmetric atrous pyramid module (AC-ASPP) utilizing convolution kernels of different scales is added at the end of the downsampling, which reduced the computational complexity and improved the computational efficiency of the model while keeping the receptive field unchanged. The effectiveness of the model is verified using a dataset of CT scan images of coal-rock combinations. The research results indicate that the VRA-UNet model performs well in crack extraction and recognition, with an average intersection to union ratio, pixel average value, and recognition accuracy of 85.22%, 90.80%, and 91.95%, respectively; Compared with mainstream segmentation networks UNet, PSPNet, DeeplabV3+, FCN, and SegNet the average intersection to union ratio of the VRA-UNet model has increased by 6.05%, 16.7%, 10.77%, 6.87%, and 6.4% respectively. The average pixel value has increased by 7.13%, 13.29%, 12.84%, 7.4%, and 7.53% and the recognition accuracy has risen by 3.82%, 14.45%, 7.4%, 5.58%, and 4.31% respectively; The fractal dimension of the fracture structure identified by VRA-UNet maintains good consistency with the fractal dimension of the original CT scan fracture structure, accurately reproducing the distribution characteristics of the internal fracture structure of the coal-rock combinations.

The inherent instability of deep coal rock masses and coal-rock composite bearing structures significantly contributes to coal rock dynamic disasters in mines. An experimental investigation into the dynamic response behavior of composite coal-rock was carried out by conducting impact loading experiments using a separated Hopkinson pressure bar system with a 50 mm diameter. The study analyzed the effects of factors like strain rate (A: 50–350 s⁻¹), lateral confining pressure (B: 4–16 MPa), and compression ratio (C: 1–4). Subsequently, a response surface experimental design with three factors and levels was developed using Design Expert12 software to analyze the impacts of individual factors and factor interactions on dynamic compressive strength σ and energy consumption rate k. Finally, a superposition damage constitutive model was formulated based on principles such as superposition theory and Weibull distribution. The results indicate that the mechanical properties of coal rock are notably affected by both individual factors and the interaction between factors. Specifically, the influence of single factors on σ and k follows the order A>C>B, while the impact of factor interactions on σ is ranked as AC>AB>BC, and on k as AB>AC>BC. The coal components within composite coal rock primarily undergo shear failure, leading to the development of macroscopic cracks in regions with lower ratios of rock components. Furthermore, the strength distribution of ternary composite coal rock demonstrates regional characteristics. The strength hierarchy of composite coal rock, from lowest to highest, follows this sequence: coal component non-interface area, coal component interface area, low component rock interface area, low component rock non-interface area, high component rock interface area, and high component rock non-interface area. The developed damage constitutive model effectively portrays the dynamic response relationship of composite coal rock, demonstrating a fitting coefficient of at least 0.97 between the theoretical and experimental curves. These research findings offer valuable insights for investigating the phenomenon of dynamic and static load superposition in the mining face during operations and for disaster prevention and control.

In the process of underground coal gasification and geothermal energy mining, the change of high temperature environment and stress field will affect the mechanical properties and strength characteristics of rocks, and the study of the change of the strength and mechanical properties of the rocks after high-temperature treatment under the action of different triaxial stresses has a very important role in the practical application of underground engineering. Based on the dynamic monitoring test system of hydraulic wetting range of coal rock under true triaxial stress, the loading and unloading tests of sandstone under different temperatures and different triaxial stress conditions were carried out. The deformation characteristics, strength characteristics and energy variation law of sandstone after high temperature were analyzed, and the effects of temperature and triaxial stress on the macroscopic strength of sandstone were studied. The failure of sandstone is dominated by brittle failure. The failure surfaces of sandstone at different temperatures are formed along the direction perpendicular to the minimum principal stress. The fracture angle increases with the increase of temperature and tends to be vertical. The bearing capacity of sandstone increases first and then decreases with the increase of temperature. The bearing capacity reaches its maximum at 600 ℃ and its minimum at 1 000 ℃. The influence of temperature on the elastic modulus, deformation modulus and other deformation parameters of sandstone is obvious, and the influence of triaxial stress on the deformation parameters is relatively small. The deformation parameters of sandstone are negatively correlated with temperature, and the deformation parameters of sandstone are positively correlated with triaxial stress. At 1 000 ℃, the elastic modulus, deformation modulus and peak strength of sandstone are the smallest, and the ultimate strain is the largest. During the loading and unloading process of sandstone, the change of energy density is roughly the same as that of peak strength, and the higher proportion of elastic energy to input energy has a greater impact on sandstone failure. Combined with the changes of mineral composition, pore fissure structure and porosity of sandstone, it is found that the macroscopic strength change of sandstone is consistent with the microscopic structure change.

The strength of coal, mudstone, sandstone and other common rocks in the surrounding rock of deep roadway is low and the fractures are developed. Dynamic disasters are prone to occur in the complex mechanical environment of deep roadway. The key to clarify the occurrence of dynamic disasters is to study the dynamic response and energy dissipation law of the internal fracture characteristics to the rock mass. In order to study the influence of internal fracture dip angle and fracture distribution on the dynamic mechanical response and energy dissipation law of rock-like mass, the rock-like mass samples with internal fractures were prepared by sand powder 3D printing. The dynamic compression test of the samples was carried out by split Hopkinson pressure bar (SHPB). The high-speed camera was used to observe the crack development, and the dynamic failure characteristics were analyzed by combining the energy dissipation principle and fractal theory. The results show that with the increase of the dip angle of the internal fracture, the dynamic peak stress and the dynamic elastic modulus of the samples decrease first and then increase, both of which are the lowest at 30º. With the increase of the number of fractures, the influence of vertical distribution fractures on the dynamic peak stress of the sample is greater than that of horizontal distribution fractures. The energy dissipation of sample failure decreases first and then increases with the increase of fracture dip angle. When the fracture dip angle is small, the existence of loose filling will aggravate the attenuation of stress wave and increase the energy dissipation. With the increase of fracture dip angle, the influence of filling on the energy dissipation of sample decreases gradually. Under the same crack inclination angle, with the increase of fracture number, the energy dissipation of vertical distribution fracture samples are less than that of horizontal distribution fracture, which is positively correlated with the dynamic peak stress of sample. With the increase of fracture dip angle, the fractal dimension of the sample increases first and then decreases. When the fracture dip angle is 30°, the fracture degree of the sample is the largest, and the fractal dimension of the sample is the largest. The increase of the number of cracks makes the fragmentation of the sample gradually uniform, and the fractal dimension of the sample shows an increasing trend. The research results reveal the influence of internal fractures on the dynamic characteristics and energy dissipation of rock mass, and provide an important basis for the dynamic test of complex internal fractured rock mass by using sand powder 3D printing technology. It is of great significance to further understand the fracture and instability mechanism of surrounding rock in deep soft rock roadway in practical engineering.

Methane is an unconventional natural gas in coal seams. Coalbed methane (CBM) migration behavior constitutes a core issue in coalbed methane extraction. However, there is no consensus on the gas migration mechanism of dual-porosity coal seams. To elucidate the mechanism of gas migration and visualize the dynamic migration process of gas, two dual-porosity borehole gas seepage models were established. Specifically, the pressure gradient drives gas in the fracture, and gas in the coal matrix is driven by the pressure gradient and density gradient respectively. The finite difference method is employed to solve both models. Through the self-developed numerical simulation software, the gas pressure distribution, gas emission velocity, and gas accumulation emission amount were obtained. By comparing the numerical results with the field-measured data, the accuracy and disparities between the two models were investigated and discussed. The results show that: ① In the initial stage of extraction, both models are predominantly governed by Darcy flow, and their gas emission velocity and cumulative gas emission quantity are substantially in consonance with the field data. In the subsequent stage, the gas within the coal matrix assumes a dominant role in gas migration. ② Due to the existence of free gas, the variation range of gas pressure in borehole in the density gradient model is larger than that in the pressure gradient model. The diffusion behavior in coal matrix is more consistent with the density gradient model in coal matrix. ③ The gas emission rate of the borehole exhibits a positive correlation with the original gas pressure, porosity, and fracture permeability coefficient, and a negative correlation with the matrix radius. The dual-porosity borehole gas transport model driven by the gas pressure gradient in fracture and free gas density gradient in coal matrix can reflect the physical behavior of borehole gas transport in coal reservoir more truly and accurately.

In order to grasp the thickness ratio effect of soft and hard interbedded overburden structure on the cross–fusion of pressure relief gas transportation and storage area, the soft and hard interbedded overburden structure with different thickness ratios was taken as the test object, and the two–dimensional physical similarity simulation experiment platform was used to carry out the crack evolution and breaking mechanism experiment of soft and hard interbedded overburden structure. Combined with the fractal theory, the fracture distribution characteristics of soft and hard interbedded overburden structure were quantitatively described. According to the theory of mining overburden elliptic paraboloid zone, the characteristic parameters such as rotation angle, penetration degree and fracture rate were introduced to study the dynamic change characteristics of cross–fusion of gas migration area and reservoir area in soft and hard interbedded overburden structure affected by the thickness ratio of soft and hard interbedded structure.The experimental results show that the internal and external boundaries of the gas transport–storage area and the transport–storage boundary are located in the mutation areas of fractal dimension, fracture penetration, fracture rotation angle and fracture rate, respectively. The fracture rotation angle of the migration area is greater than 2.07°, and the fracture rotation angle of the reservoir area is 1°～2.07°. The fracture connectivity of gas migration area and reservoir area is 0.6～1.0 and 0.2～0.6, respectively. According to the changes of fracture rate, rotation angle and penetration degree in gas transportation–reservoir area, the boundary of gas migration area, reservoir area and cross fusion area is determined. The fracture rate, rotation angle and penetration degree decrease rapidly at first, then the decrease rate slows down obviously, and finally the decrease rate increases again. The overlying strata of the soft and hard interbedded strata experienced five periods: the formation of the transport–reservoir area, the first appearance of the migration area and the reservoir area, the first formation of the cross–fusion area, the expansion of the cross–fusion area, and the gradual blurring of the transport–reservoir area and the cross–fusion boundary. Finally, a complete elliptical parabolic banded overburden fracture field was formed. The quantitative characterization model of thickness ratio effect of gas transport–storage area in soft and hard interbedded overburden structure is established. At the same time, the boundary and state determination process are determined according to the characterization parameters of gas transport–storage area, and the fracture evolution and fracture mechanism of overlying strata in soft and hard interbedded overburden structure are determined.

Pulsating hydraulic fracturing technology has demonstrated advantages such as lower initiation pressure and complex fracture networks in reservoir stimulation fields including gas extraction, shale gas development, and geothermal exploitation. Currently, the method of controlling flow rate to achieve pulsating hydraulic load output is commonly used in engineering, but the mechanism of fracture propagation remains unclear. To address this, true triaxial flow-controlled pulsating hydraulic fracturing experiments were conducted on sandstone specimens to analyze the effects of pumping frequency and rate on injection pressure, acoustic emission energy, fracture type, and macroscopic fracture morphology. The results indicate that the initiation pressure, breakdown pressure, and transient energy of flow-controlled pulsating hydraulic fracturing increase with the increase of pumping frequency and rate, while the fracture propagation area decreases with increasing pumping frequency and initially increases then decreases with increasing pumping rate. Compared to conventional hydraulic fracturing, the acoustic emission energy is more concentrated, the initiation pressure is lower, with a maximum reduction of 21.9%, the proportion of shear cracks increases by 5.6% to 17.8%, and the fracture propagation area can increase up to 2.3 times. The pulsating hydraulic load varies complexly, showing a ramp-shaped load with fixed frequency, increasing pressure mean and amplitude during the initiation phase, and a horizontal cyclic load with fixed upper pressure limit, frequency, and amplitude during fracture propagation. The upper, lower, and mean pressure limits are positively correlated with pumping frequency and rate, while the pressure amplitude is negatively correlated with pumping frequency and initially increases then decreases with increasing pumping rate. Different types of output pulsating hydraulic loads result in significantly different fracture propagation characteristics. To enhance fracturing effectiveness, the design of fracturing parameters must ensure that the output pulsating hydraulic load has sufficient strength and amplitude.

In order to clarify the effect and mechanism of supercritical CO₂ pulsation on low-rank coal, deep low-rank coal from the Aiwiergou Mine in Xinjiang was collected as the experimental object. Experimental research was carried out through an independently built supercritical CO₂ pulsation fracturing experimental system. Before and after the pulsation, the coal samples were respectively measured by XRD, FTIR, low-temperature liquid nitrogen adsorption test, low-field nuclear magnetic resonance, and uniaxial compression test. The changes in the microscopic structure, pore structure, and mechanical properties of the coal before and after the supercritical CO₂ pulsation were quantitatively analyzed. The results show that after the supercritical CO₂ pulsation, the mass fractions of calcite and clay minerals in the coal increase. The change in the mass fraction of clay minerals is mainly affected by the increase of ammoniated illite. The peak values of various organic functional groups in the coal show different degrees of decrease, indicating that the extraction effect of supercritical CO₂ on organic matter is obvious. After the pulsation, the microcrystalline structure of the coal is damaged, the crystal lamellar spacing increases by 0.002 9 nm, and the coal structure tends to be loose. Compared with before the supercritical CO₂ pulsation, the effective porosities of coal samples XJ-A and XJ-B increase by 66.7% and 128.6% respectively after the pulsation. The connectivity of pores inside the coal is enhanced, and the pulsation has a more obvious effect on the modification of micropores and mesopores. After the supercritical CO₂ pulsation, the mechanical properties of the coal deteriorate significantly. The uniaxial compressive strength and elastic modulus decrease by 54.73% and 59.82% respectively. The acoustic emission ring-down count and cumulative energy are both greatly reduced. The peak value of the ring-down count drops to 51% of that before the pulsation, the cohesion of internal particles in the coal sample decreases, and the cumulative energy decreases by 62%. In conclusion, the supercritical CO₂ pulsation can significantly damage the micro-macro structure of low-rank coal, which is of great significance for the fracturing and permeability enhancement of deep low-rank coal reservoirs and the extraction of coalbed methane.

There are many practical problems in the spontaneous combustion area of small coal mines coal mine in Shanxi Province, such as many fire sources and wide distribution range, hidden fire source location, difficult to accurately detect, many air leakage channels, and difficult to prevent and control. In order to accurately detect the location and range of fire source, firstly, The detection mechanism of spontaneous combustion area detection method is studied. Based on the basic theory of radon measurement method and magnetic method to detect fire area, the detection principle of ground “radon measurement-magnetic” joint technology is proposed to accurately detect the location and range of small kiln fire source. Secondly, FD213 alpha spectrum radon measuring instrument and GSM-19 T magnetometer were used to collect the hidden fire source data of small kiln on site. The ground “radon measurement-magnetic” joint technology was used to process and analyze the collected data, and the location and scope of the hidden fire source of small kiln were comprehensively delineated. The location of the fire source was verified by using underground temperature measurement and gas measurement boreholes, combined with the preliminary detection results and the distribution of underground roadways. Finally, the borehole grouting technology is used to control the fire area, and the combined technology of beam tube gas monitoring and re-measurement of radon method is used to detect the effect of fire area control. The test results show that the location and range of the hidden fire source delineated by the “radon-magnetic” joint technology are successfully verified by the results of downhole temperature measurement and gas measurement drilling. The area of the hidden fire area delineated in this test is about 213 m². The cumulative filling material of drilling grouting fire area treatment is about 1862 m³; the detection of the treatment effect shows that the internal index gas CO concentration in the fire area is delineated.

The method of sensing coal spontaneous combustion temperature based on acoustic wave technology has the advantages of wide application scenarios and real-time continuity, but the precursor characteristics of acoustic wave signals in the process of coal spontaneous combustion have not yet been revealed. The study of acoustic effects and precursor characteristics in the process of coal spontaneous combustion provides a theoretical basis for the detection and warning of coal spontaneous combustion by acoustic wave method. The article firstly theoretically analyses the thermal damage rupture evolution process and acoustic wave transmission characteristics of coal rock. Three sizes of coal spontaneous combustion acoustic wave information testing systems were established to test the infrasound, acoustic emission and acoustic sound velocity information during coal heating and combustion. Using linear fitting, multiple fractal theory, Fourier transform and other methods, the paper analyse the temporal characteristics, temperature correlation, spatial characteristics, nonlinear characteristics and spectra of acoustic signals to reveal the precursor characteristics of acoustic signals in the process of spontaneous combustion of coal. The results show that with the increase of coal temperature, the infrasound sound pressure value increases paroxysmally, and the ringing counts, energy value and acoustic sound velocity of acoustic emission signals are positively correlated with the temperature. The infrasound signals and acoustic time changes were more obvious in the pre-coal spontaneous combustion period. The acoustic emission signal increases and changes more significantly after 100 ℃. The acoustic signal has spatial characteristics. With the increase of the distance from the heat source, the ringing count and energy change by decreasing. The speed of sound at different distances increases with increasing temperature, and the R² of the speed-temperature fitting equation exceeds 0.9. The infrasound and acoustic emission signals have multiple fractal characteristics during coal heating and correspond well with the thermal rupture of spontaneous coal combustion. In addition, the main frequency of infrasound waves migrated and the amplitude of the main frequency increased during the process of coal warming. Finally, the acoustic emission test was carried out in the high-temperature anomaly area of the coal field. The acoustic emission signals in the temperature anomalies change significantly.

Three-dimensional geological modeling (3DGM) in mining offers an intuitive representation of underground geological structures. This visualization significantly enhances the precision and efficiency of mining operations. Additionally, it helps to reduce risks and optimize resource utilization. This paper summarizes the principles and methods of existing 3DGM technologies. It reviews the current application status of 3DGM in mining development. Furthermore, it discusses the challenges faced by current 3DGM technologies and their future development trends. 3DGM is to use computer 3D graphics technology to integrate scattered exploration data into a digital model reflecting geological structure and property changes. These models can be divided into surface model, volume model and mixed model. According to the purpose and data source, it can also be subdivided into static model (structural model, stratigraphic model, attribute model) and dynamic model. The modeling process includes data acquisition, processing and model construction. It involves diverse data sources such as geological exploration data, remote sensing data and engineering data. A variety of interpolation methods are used to make up for the data deficiency. Various modeling software has been developed both domestically and internationally. In mining applications, 3DGM is primarily used in mineral resource development, mine design, and safety management. By building accurate 3D models, it improves the understanding of underground structures and enables intelligent resource estimation and dynamic management. 3D ventilation simulation and visualization technologies optimize ventilation systems and reduce costs. Furthermore, integrating 3DGM with automated equipment and sensor technologies facilitates real-time monitoring, precise positioning, and dynamic adjustments in mines. In mine disaster monitoring, 3DGM enables the graded assessment of mining hazards. Combined with numerical simulation techniques, it provides effective support for disaster prevention and risk mitigation, ensuring the safety of personnel. However, current 3DGM technologies face challenges such as low data processing efficiency and insufficient automation. Future developments should integrate artificial intelligence to advance 3DGM towards greater intelligence and automation.

Specialized drilling technology is a critical component of drilling technology system in underground coal mine. Its core lies in adopting specialized drilling method and equipment to address unique working conditions and achieve specific drilling objectives, playing a vital role in preventing major geological disasters and advancing geological transparency in underground coal mine. The typical application scenarios of specialized drilling technology are summarized, such as precise gas control in soft-fragmented coal seam, replacing tunnel with borehole for relieved gas drainage in roof, mine water detection and release, and grouting reinforcement of water-resisting layer.It delves into the key technical challenges encountered during implementation.Focusing on drilling equipment upgrade, innovative drilling method, and drilling parameter optimization, the study comprehensively reviews the current development status, critical technological advancements, and practical outcomes.The specialized drilling technology suitable for soft-fragmented coal seam, hard rock strata, and fractured and water sensitive strata are outlined, and the existing technical bottlenecks and challenges are identified.Given the evolving trends of diversified demand, equipment intelligence, and region-specific adaptation in underground specialized drilling technology, future research priorities are proposed: ① To enhance air directional drilling technology for soft-fragmented coal seam to improve formation adaptability and process compatibility. ② To advance high-efficiency drilling acceleration technology and tool for hard rock strata to boost rock-breaking efficiency and service life of impact drilling tool. ③ To overcome theoretical and technical barrier in pressure-controlled drilling under near-horizontal condition to strengthen borehole formation capability in fractured and water-sensitive strata of roof or floor. ④ To accelerate the improvement of the reliability and stability of intelligent drilling technology and equipment in underground coal mine, and promote the realization of the transformation of specialized drilling technology from equipment intelligence to engineering intelligence. ⑤ To research and constructe specialized drilling technology system of regional, mining area and mine, and promote their wide application and in-depth development in the coal mining industry.The study can provide an important reference for the technological innovation and efficiency improvement of drilling process in underground coal mine.

The deep coal reservoir in Qinshui Basin has the geological characteristics of low permeability to ultra-low permeability, local high saturation, low porosity and high thermal evolution degree. Under the development mode of vertical well/horizontal well + large-scale fracturing, it is necessary to adapt the completion technology to ensure the reservoir reconstruction and late production effect. This paper systematically analyzes the different well structure and tool types of conventional single casing completion technology, open hole/screen completion technology, casing + slip-sleeve completion technology, and discusses the applicable conditions and technical characteristics of different completion technologies. The results show that the conventional single-casing completion technology is mature, and the three-way completion structure is widely used, but it is not conducive to cost reduction and efficiency increase. The second well completion structure is only applicable to the geological conditions where the overlying strata is relatively stable and there is no aquifer or weak aquifer. Open hole/screen completion technology can reduce the pollution of coal section, but it can not carry out late reservoir transformation, and is suitable for the reservoir with complete coal structure and strong permeability of coal seam. Casing + slip-sleeve completion can eliminate perforation before fracturing, but the normal operation of downhole tools is not guaranteed and the fracturing rate is limited in the later stage. For producing vertical Wells in U-shaped development Wells, a new double-casing completion technology was proposed on the premise that the wellbore could not meet large-scale fracturing operations. The principle of this technology is to reshape the wellbore by selecting casing and cementing tools according to geological conditions and characteristics of the old wellbore, so as to meet the late fracturing process and ensure the fracturing reconstruction effect. It is mainly characterized by the use of cement sealing above the roof of the target coal seam, with high pressure resistance and large inner diameter casing, so that the wellbore can meet the large-scale fracturing of high construction pressure. At the same time, the selection of screen pipe in the coal seam saves the perforation cost, and can effectively avoid the formation collapse and bury the coal seam in the drainage stage. The technology has been successfully applied in deep coal reservoirs in Qinshui basin, and the test results show that it can meet the needs of deep coalbed methane fracturing operations under large scale and high construction pressure, and has the prospect of popularization.

The scarce coking coal is the fundamental guarantee energy of steel and chemical industry, and is considered as a strategic mineral resource because of its scarcity, irreplaceability and non-renewability. The Huaibei Coalfield in Anhui Province, as one of the first key mining areas of scarce coking coal set up by the state, has rich resources and excellent coal quality, and is the most promising coking coal resource source in East China.In order to systematically understand the occurrence characteristics and exploitation and utilization potential of scarce coking coal in the study area, this paper reveals the spatio-temporal difference characteristics, formation control factors and resource potential of scarce coking coal in the study area through coal rock analysis, industrial coal quality index test, comprehensive mapping and resource evaluation, using previous drilling data and underground supplementary sampling tests.The results show that the scarce coking coal in Huaibei coalfield is well developed, and the floating coal after washing shows the characteristics of ultra−low water, ultra−low ash, ultra−low sulfur, ultra−low fixed carbon, medium and high volatile, and has the process properties of medium−strong bonding and coking, low potassium and sodium content, and good mechanical strength and thermal strength.The scarce coking coal is controlled by the regional structure and the sedimentary distribution of coal seams, and the coal seams of Xiashihezi Formation 7 and 8 are the most developed. The scarce coking coal is distributed sporadically in the northern part of the coalfield, and continuously distributed in the southern part of the coalfield, and the metamorphism degree of coal is increasing from southwest to northeast in the area.The formation and spatiotemporal distribution of scarce coking coal were controlled by the fault activity in different directions and magmatic coupling in the post−Yanshanian period. The scarce coking coal is divided into three grades: Low−ash−ultra−low−sulfur coal, ultra−low−ash−low−sulfur coal, low−ash−low−sulfur coal, and low−ash−low−sulfur coal from the perspective of clean and green development.It is estimated by classification, classification and zoning that the reserved resources of shallow scarce coking coal at −1500 m are about 10.165 million tons, accounting for 78.13% of the total reserved resources in the coal field, and the resource potential is huge. By establishing an evaluation index system for the development and utilization of scarce coking coal resources, and using the hierarchical fuzzy evaluation method, each mine and exploration area are divided into four types of development zones: favorable, relatively favorable, general, and unfavorable. Suggestions for protective and restrictive development, total quantity control development, and prohibitive development are proposed for different types of development zones.

Under the background of comprehensive utilization and carbon neutralization of coal-series associated minerals, coal-series graphite has attracted more and more attention from scholars at home and abroad. As one of the important characterization means to study the degree of molecular order and determine the structure of coal microcrystals, XRD spectrum analysis is of great significance to reveal the relationship between the order of carbon microcrystal arrangement and the graphitization degree of coal. In order to obtain the XRD related index parameters of coal more scientifically, quickly and effectively, and constantly improve the evaluation index system of coal-series graphite, taking 7 pairs of metamorphic coals in Huaibei coalfield as the research object, with the help of coal quality analysis, vitrinite reflectance and XRD, the effects of different peak fitting methods and different sample pretreatment methods on XRD spectral parameters were studied, and the effective determination method and response law of coal-series graphite evaluation index based on XRD spectral analysis were discussed. Research shows that; ① Different peak fitting methods have great influence on the evaluation results of coal-series graphite, and whether the bottom back is deducted has more obvious influence on the experimental results than whether the baseline is removed; ② Whether the (γ) peak is considered has a great influence on the experimental results, which directly affects the classification and identification results of the graphitization degree of the samples; ③ By analyzing the XRD spectrum curves of different raw coal and standard Si powder ratios, it is concluded that under the mass ratio of raw coal/SiO₂ ≈3∶1, the 2θ value of (002) peak of raw coal sample can be effectively determined by internal standard method, and the experimental effect is equivalent to that of demineralization method. ④ V_daf and H/C can be recommended as the main coal quality parameters, while La, La/Lc and d₀₀₂ can be recommended as the main structural parameters. Among them, V_daf < 20%, H/C > 0.035, La < 4 nm, La/Lc < 1 and d₀₀₂<0.344 nm can be used as the critical marks for the coal metamorphism in Huaibei coalfield to enter the semi-graphitization stage. The above research results provide a reference for the optimization of evaluation index and the construction of identification system of coal-series graphite in Huaibei coalfield.

The development height of water-conducting fracture zone in thick weakly cemented strata is of great significance to coal production safety and groundwater resource protection in Shaanxi-Inner Mongolia weakly cemented mining area. The study found that the overlying Cretaceous rock layer in the thick weakly cemented mining area has the characteristics of large thickness, low strength, poor cementation and undeveloped joints. However, during coal mining in this area, the overburden is severely damaged and the water-conducting fracture zone is highly developed, with a fracture-to-mining ratio of about 30, seriously threatening the production safety of the mine. How to accurately predict the development height of water-conducting fracture zone in thick weakly cemented overburden has become one of the keys to safe and efficient coal mining in thick weakly cemented mining area. Based on this, this paper takes a working face in Shilawusu mining area as the research background, and uses UDEC numerical simulation software combined with measured data to explore the evolution characteristics of water-conducting fracture zone in thick weakly cemented overburden. It is concluded that the overall shape of water-conducting fracture zone in thick weakly cemented overburden is “△”, which is significantly different from the "saddle" characteristics in the east. Moreover, its shape and development height are affected by mining height, thickness and location of weakly cemented rock layer and other factors. The thick weakly cemented overburden has a significant control effect on the development shape and height of water-conducting fracture zone. On this basis, a prediction model for the development height of water-conducting fracture zone in thick weakly cemented overburden is established using the plate theory, and it is applied in the Shilawusu mining area, verifying the effectiveness and accuracy of the established model.

With the continuous development of China’s coal resources to the deep end, the complex conditions of rock stratum existence bring great challenges to the construction and use of shafts. Deep vertical shafts often traverse multi-layer aquifers, and the seepage problem of the shaft wall caused by roof pressure water is particularly serious, which not only threatens the mine safety, but also brings challenges to the traditional grouting management methods, such as high cost, difficult construction and repeated seepage problems. In this regard, a set of visual similar simulation experiment platform was established based on the hydrophobic pressure reduction theory, simulated the water pressure of the well wall and the change of the flow rate of the water discharge holes under the conditions of different rock formations and different number of water discharge holes, analyzed the flow rate and seepage law of the water discharge holes, the development law of the morphology of the groundwater landing funnels, and the change law of the seepage water pressure and the pressure law of the well wall. And take Binchang Wenjiapo coal mine as an example to carry out on-site water discharge test. The results show that: the flow rate at the outlet increases with the increase of pipe diameter; the water flow rate of two holes increases more than that of a single hole, and the water flow rate of three holes increases less; the closer to the well, the deeper the water level drops, the influence range of the landing funnel increases with the increase of the number of discharge holes, the shape of the landing funnel of the single-layer water discharge experiment is kept symmetrically with the descending holes, and the bottom end of the funnel shows a conical shape. The water pressure decreases quickly in the early stage of water discharge, and tends to stabilize in the later stage; the results of the simulation experiment are applied to the field, and the single-hole and double-hole water discharge tests are carried out by drilling sparse discharge holes between the main and auxiliary wells of the Wenjiapo Mine, and the stabilized water influx in the single-hole is 75 m³/h, and the stabilized water influx in the double-hole is 175 m³/h. The test has significantly lowered the groundwater level, and formed the landing funnel with a maximum depth of 40 m. The actual measurement in the field is more in agreement with the results of the simulation experiment, which verifies the similarity with the results of the simulation experiment. Simulation test results are more consistent, verifying the scientific and practical effect of the hydrophobic bucking management method. The results of this research can help to promote the progress of wellbore water seepage control technology under the influence of deep roof pressurized aquifer.

In the development of large-scale open-pit coal fields in China, there are often several Coal companies entering the same coal field, and each Coal company is a main body of development, mining coal resources in the state-allocated strip area, and building several open-pit mines at the same time or successively. Adjacent open-pit mines, during the excavation process, are susceptible to landslide hazards, particularly on the slopes where the adjacent end gangs, influenced significantly by the mining activities of both mines, reach the boundary. The Shengli West 2 open-pit coal mine and Wulantuga Zhemei open-pit mine are neighbouring surface mines in the same coal field, and the mining rights of the two mines do not overlap at the neighbouring end gangs, resulting in the formation of a trapezoidal platform between the west gang of the West 2 open-pit mine and the east gang of the Zhemei open-pit mine, and the impact of the overlying fault fragmentation zones and the operations of the two mines, which resulted in the appearance of signs of geologic hazards within the platforms, such as bottom drums, fissures, and collapse of faults. With the development of the stripping project of the two mines and the leaning operation of the adjacent end gangs, a steeper irregular platform will be generated between the adjacent end gangs and the boundary slopes, in order to avoid the occurrence of catastrophic landslides, and on the basis of the stability analysis of the steep platform, the treatment method of “digging out and backfilling as a whole” was put forward and the specific treatment plan was determined. The management plan specifies the scope of operation of the “overall excavation and backfilling” management method, and designs the shape of the clean-up work gang and the slope of the internal discharge site when the management is completed, which can completely eliminate the hidden geological hazards and release 9 318 300 m³ of internal discharge space in the West Second Open Pit Mine and 7 646 900 m³ of internal discharge space in the Zhemei open-pit mine Mine, so that the two mines can continue to achieve complete internal drainage. The results of this research provide the necessary theoretical basis and technical support for the treatment of geological hazard hidden bodies in the adjacent fault zones of the West 2 and Zhemei open-pit mines, and also provide relevant experience for the implementation of geological hazard management in similar adjacent open-pit mines in the same coal field.

Slope monitoring is the last line of defense to ensure the safety of worker and equipment in open-pit mines. A sound slope monitoring and warning system and accurate warning thresholds are important ways to achieve monitoring results. However, existing monitoring systems have problems such as independent operation, lack of complementarity, and lack of basis for setting warning thresholds. This is an important scientific problem that urgently needs to be solved in open-pit mine slope monitoring in China. Taking gravity landslides as the research object, based on the characteristics of the cumulative displacement time curve, they are classified into three types: gradual, sudden, and stable; According to the shape of the sliding surface, it can be divided into two categories: circular sliding and planar sliding. We compared and analyzed the advantages and disadvantages of slope radar and GNSS monitoring, and established an error calculation model for the displacement and deformation rate of slope radar monitoring units based on a rigid body motion model. Then, combining the advantages of the two monitoring methods, we proposed a slope safety evaluation slope radar coverage monitoring GNSS accurate warning system for open-pit mines. Based on the tangent angle warning criterion, a four level warning threshold for slope deformation rate was proposed using the critical tangent angle tangent value multiple method. Furthermore, the critical tangent angle was determined using the T-value equal division method during the acceleration stage of similar landslides. Two key tangent angles and warning threshold determination methods are proposed for open-pit mine slopes that have not experienced landslides. A study was conducted on two open-pit mine slope landslide cases, determining the uniform velocity stage and critical tangent angle, and providing a four level warning threshold. The threshold was applied to slope deformation warning, and the results showed that the threshold of open-pit mine No.1 can timely and accurately determine the warning level of the eastern slope deformation, and reserve 6 hours of emergency evacuation time after the first level I warning information is issued; The key tangent angle of open-pit mine No.2 can be applied to determine thresholds for different stages, but the average rate needs to be accurately determined in conjunction with the S-t curve of specific monitoring stages. The deformation monitoring system and warning threshold determination method provide strong support for the reliability of open-pit mine slope monitoring.

In response to the technical challenges of achieving intelligent comprehensive management of the entire lifecycle of coal mining equipment, the development process of China’s coal mining equipment lifecycle management technology is divided into three stages. The pain points in China’s current coal mining equipment management are analyzed from the aspects of standard system construction, equipment ledger process management, intelligent operation and maintenance management, and system platform construction. We have studied the management objectives, business processes, and implementation entities of coal mining equipment in the early, middle, and later stages, and proposed the overall architecture of a digitalized management system for the entire lifecycle of coal mining equipment. The main influencing factors of coal mining equipment selection optimization design were analyzed, and a calculation method for comprehensive mining equipment selection optimization was studied. An optimization algorithm and software platform for comprehensive mining equipment selection were developed, and a method for constructing a series spectrum and coding system for coal mining equipment was proposed. Dividing coal mining equipment management into six business domains and clarifying the interrelationships between different business domains; We studied the business division of labor in different stages of the work order management process and proposed a refined work order management process based on AI empowerment. Analyzed the differences between different operation and maintenance modes, and proposed a technical framework for coal machine equipment operation and maintenance management based on the regional centralized operation and maintenance mode. Analyzed the current status and existing problems of coal mining equipment state perception technology, studied indirect perception methods based on non-contact sensing technology, and reduced installation and operation costs while improving solution accuracy. We studied the technical path of remote operation and maintenance of physical models based on digital twin technology, proposed a geometric-mechanism-data model fusion driven operation control and maintenance strategy, and looked forward to the development direction of intelligent comprehensive management technology and system platform for the entire life cycle of coal mining equipment.

In view of the problems of typical unstructured, degraded characteristics and large-scale environment in coal mine tunnel, and the 3D reconstruction of coal mine tunnel is prone to low pose estimation accuracy and large cumulative drift error, a 3D reconstruction method of coal mine tunnel fused with lidar and inertial measurement unit (IMU) is proposed. In this method, the residual function of the lidar observation model and the prior state deviation of the IMU prediction model are tightly coupled through iterative kalman filter, and a more accurate posteriori state is obtained through state update, which provides reliable pose estimation for the degraded environment. In order to reduce the cumulative drift error in the process of tunnel 3D model reconstruction, a loopback detection algorithm based on Voxelized Generalized ICP (VGICP) is proposed. It is registered in a voxel-based single to multi-distribution mode, so as to complete the selection and accurate matching of loopback keyframes, realize the global pose correction of loopback keyframes, and effectively reduce the cumulative drift error of 3D reconstruction of coal mine tunnel. Compared with the A-LOAM and LEGO-LOAM algorithms, the proposed algorithm has significantly improved the accuracy and global consistency of pose estimation. Experimental results on public datasets show that the root mean square errors of RPE and APE of the proposed algorithm are 0.2718 and 0.5008, respectively, which are reduced by 53.14%, 50.97%, 48.31%, and 50.41%, 47.99% 47.49% respectively compared with other algorithms. Two experiments of simulated coal mine tunnel are carried out. The results show that the error percentage of each area of the indoor corridor model constructed by the proposed algorithm in the directions of length, width and height was within 1.2%. The three-dimensional model of coal mine tunnel constructed by the proposed algorithm is consistent with the spatial distribution of the real coal mine tunnel. The overall distance degradation error is only 2.46%, which is 66.12%, 65.30% and 70.43% higher than that of the other three algorithms, respectively. Three-dimensional reconstruction experiments are carried out in the roadway excavation in the main laboratory of the coal mine. The results show that the average error percentages in the length, width and height directions are 0.47%, 0.75% and 0.67%, respectively. It can realize the accurate three-dimensional modeling of the tunnel.

Aiming at severe working conditions of fully mechanized mining face, poor passage, weak protective ability, and limited means of perception technology exists for intrinsic safety inspection robot and cann't be normally operated. The explosion proof inspection robot scheme is proposed for severe working conditions. And three characteristic working conditions: climbing the hill, shear, torsion are analysized for railed inspection robot of mechanized mining face. Based on three kinds of harsh working conditions, a kind of multiple degrees of freedom explosion-proof inspection robot is designed. Besides the video perception method, the lidar slam mapping ability is added, then, the precise 3D perception capability is provided. And the perception is not affected by lighting conditions, so it is can run in lightless conditions. For climbing the hill working condition, tracked walking mechanism is adopted. For shear working condition, a kind of walking mechanism which has front and rear differential drives are designed and can steer on its own. For torsion working condition, longitudinal rotated axis is added. In order to control this inspection robot effectively and ensure steady running in working face, the kinematics model is analysized and gotten aiming at railed mechanized mining face inspection robot, and is subjected to trajectory and chassis constraints. The control method of the inspection is introduced. The reference trajectory is extracted from the point cloud which is obtained from slam. The velocity of every tracked active wheel is calculated according to the reference trajectory. A kind of MPC control method is proposed based on the kinematics model, and the MPC is implemented only for front differential drive, the control setpoint of rear drive is calculated by the speed reference of front differential drive and the constraint between them. This method balances between the computational complexity and real time performance. And the algorithm is realized based on ROS. Lastly, through enough underground industrial tests, this method is effective for climbing the hill, shear, torsion working condition and inspection tasks can be executed by normalization.

Large scale open-pit mining activities have had a significant impact on the topography, geomorphology, and ecological environment of the loess gully region. In order to alleviate the common problems of insufficient internal drainage space in open-pit mines, such as excessive slope angles, over elevation soil discharge, limited production capacity, as well as soil erosion and difficult treatment and utilization of fragmented gully land in the natural gully area around the coal mine, this paper proposes a collaborative mining and treatment technology model for ditch filling and land reclamation in the vicinity of open-pit mines. The principle is based on the analysis of technical, economic and environmental benefits. The suitable stripped rock and soil materials that were originally intended to be dumped in the dumping site are backfilled to adjacent gullies outside the mining area that are suitable for treatment, and finally covered with soil and vegetation construction, in order to save dumping space and reuse land in the gully area. The main technical steps include: extracting the location of the gully to be treated based on the algorithm of constructing concentric rectangular windows inside and outside, optimizing the earthwork allocation path of the waste dump based on the “source sink” theory, backfilling the gully area based on the reshaping of the near natural landform, screening the waste materials and reconstructing the soil layer profile of the gully backfilling, greening and land reuse of the covering soil, and evaluating the ecological effects of collaborative mining and treatment. This article takes a certain open-pit coal mine in Zhungeer Banner, Inner Mongolia as an example to conduct research. The results show that the total area of natural gullies around the open-pit mine treated by this model is 2.588 × 105 m², reducing the amount of earthwork discharge in the mining area by about 1.536 × 107 m³; By combining the CLiDE geomorphic evolution and soil erosion assessment model, we simulated and evaluated a 66.53% reduction in regional soil erosion after collaborative treatment of surrounding gullies over a 30-year period. At the same time, the landscape after collaborative treatment showed a relatively stable evolution trend, and a large amount of available land was added, achieving the goal of “one soil, two uses”. To ensure the promotion of this technological model, corresponding policy support and strict process supervision are also needed to ensure that the model generates maximum comprehensive benefits and avoids causing new land occupation and ecological environment problems.

As the most significant and representative form of mining damage in the coal mining area of northern Shaanxi in the middle reaches of the Yellow River, the soil erosion effect caused by mining ground fissures can not be ignored. In order to study the influence of mining-induced ground fissures on the anti-scourability of surrounding soil, the mining-induced ground fissures with widths of 0−10 cm, 10−20 cm and 20−30 cm in the mining-induced ground fissure development area in the north wing of Ningtiaota in northern Shaanxi were selected as the research objects. The surface soil with a horizontal distance of 0−80 cm was collected, and the soil mechanical composition, >0.25 mm water-stable aggregates and organic matter were measured. Five influencing factors of soil anti-scourability and soil anti-scourability coefficient, which are generally concerned by domestic and foreign scholars, reveal the changes of soil physical and chemical properties and anti-scourability coefficient around mining-induced ground fissures. The prediction model of surface soil anti-scour coefficient in the mining ground fissure development area of northern Shaanxi was constructed. The results show that: The sediment loss caused by the erosion of the surface soil around the mining ground fissure will show a three-stage change process of “rapid increase (0−1 min), slow increase (1−3 min) and stable (3−10 min)” with the increase of the erosion time. The sediment loss caused by the erosion mainly comes from within 3 min, and its contribution rate exceeds 90%. Mining-induced ground fissures will have a significant effect on reducing the anti-scourability of the surrounding surface soil, with a maximum decrease of 32.1%, and the greater the width of the mining-induced ground fissures, the closer to the ground fissures, the more obvious the effect. The soil anti-scourability coefficient was significantly negatively correlated with the mass fraction of sand (p<0.01), and was significantly positively correlated with the mass fraction of clay, >0.25 mm water-stable aggregates and organic matter (p<0.01). Among them, the mass fraction of sand, organic matter and >0.25 mm water-stable aggregates can be used as the main controlling factors of soil anti-scourability in the mining ground fissure development area of northern Shaanxi. The prediction model of surface soil anti-scourability coefficient in mining ground fissure development area is constructed by taking the width of mining ground fissure and the horizontal distance from the fissure as variables. It is found that when the horizontal distance exceeds 158 cm, the influence of mining ground fissure on the surrounding surface soil anti-scourability basically disappears, which can be used as the target area for the prevention and control of soil water erosion in the mining ground fissure development area of northern Shaanxi. The research results can provide a scientific basis for the precise prevention and control of soil erosion in the mining area of northern Shaanxi in the middle reaches of the Yellow River.

The large number and concentrated distribution of pollution sources in coal-related industry agglomeration area can easily induce groundwater pollution, and it is of great significance to carry out the risk assessment of groundwater pollution to protect the groundwater environment in such areas. Taking a coal-related industrial agglomeration area in Shanxi Province as study area, the DRASTIC and PLEIK models were used to evaluate the vulnerability of pore water aquifers and karst water aquifers in the study area, respectively. Analytic Hierarchy Process (AHP) was used to determine the weights of the indicators in the PLEIK model, and the integrated loading of pollutants and the vulnerability of groundwater were used to characterize the risk of groundwater pollution in the study area. At the same time, combined with the water quality level of groundwater sampling points in the study area, the Random Forest (RF) classification algorithm was utilized to construct the groundwater pollution risk classification prediction method, and the evaluation results were compared with those of the superposition index method. The results show that: ① the loading of pollution sources in the study area is high, and the proportion of high pollution source loading area is about 26.73%, which is related to the characteristics of the more concentrated distribution of pollution sources in coal-related industry agglomeration area, and when quantifying the loading of pollution sources, the superposition effect of multiple pollution sources is obvious; ② the comprehensive vulnerability of the groundwater in the study area is dominated by the medium grade, and the proportion of medium vulnerability area is about 82.59%, with the pore-water aquifer high vulnerability area The high vulnerability zone of pore water aquifer is mainly located in the eastern and southeastern part of the study area, and the high vulnerability zone of karst water aquifer is mainly distributed in the karst exposed area north of Fenhe River; ③ the areas of low, medium and high risk zones of groundwater in the study area based on the superposition index method are 3.55%, 59.67% and 36.77%, respectively, and the consistency rate with the water quality level of the actual sampling points is 75%, and the risk of groundwater pollution predicted by the use of RF The predicted risk of groundwater contamination was dominated by low risk, and the correct rate between the grading results and the water quality of the actual sampling points was 97.7%, which improved the accuracy of the results calculated by the stacked index method by about 22.7%. The evaluation results are intended to provide a basis and reference for groundwater pollution control in the study area.

Coal gangue (CG), the most abundant industrial solid waste in China, poses significant ecological and environmental risks to surrounding mining regions. Given that incorporating coal gangue into concrete is among the most efficient disposal methods, this paper provides a comprehensive review of its composition, the effects of its usage on concrete properties, and the specific conditions for using coal gangue as coarse aggregate, fine aggregate, and mineral admixture. The review emphasizes how variables such as dosage, particle size, and water-cement ratio influence the workability of fresh concrete, as well as the mechanical properties and durability- including water absorption, frost resistance, chloride ion penetration resistance, carbonation, and sulfate attack resistance-of hardened concrete. Key findings highlight significant regional variations in the composition and performance of coal gangue. When used as an aggregate, it is crucial to minimize its carbon and sulfur content, crushing value, and water absorption while optimizing dosage, particle size, and water-cement ratio. For its use as a mineral admixture, considerations include reducing particle size (to below 0.074 mm and activating the coal gangue powder.) The recommended maximum dosages for coal gangue as coarse aggregate, fine aggregate, and mineral admixture are 45%, 20%, and 10%, respectively. The paper also details performance enhancement techniques for coal gangue aggregates and mineral admixtures, such as thermal activation, surface coating with cement mortar, water glass treatment, and microbial mineralization. Thermal and mechanical activation are highlighted as key methods for boosting the reactivity of mineral admixtures. This study elucidates the mechanisms behind improving coal gangue’s performance as aggregate and admixture and identifies research gaps and challenges, offering insights for advancing coal gangue utilization and the development of sustainable building materials.

The resource utilization of coal gasification slag has become a research focus in the field of coal-based solid waste. The physical and chemical characteristics and the heavy metal migration rules of coal gasification slag are very important for its resource utilization. Firstly, three coal gasification technologies of entrained flow, fluidized-bed and fixed-bed gasification were introduced and compared. Then, the physical and chemical characteristic of coal gasification slag, including particle size composition, apparent and microscopic morphology, inorganic and crystalline mineral composition and carbon residue content, were summarized. Finally, the migration behavior, risk assessment code and leaching behavior of heavy metals in coal gasification slag were analyzed. The physical and chemical characteristics of coarse slag and fine slag are different to some extent. From the perspective of physical and chemical characteristics, the differences of fine slag and coarse slag are as follows. The particle size of fine slag is much lower than that of coarse slag, about 60% of fine slag is less than 0.250 mm, and about 50% of coarse slag is more than 0.500 mm. The specific surface area of fine slag is larger than that of coarse slag, and the average pore diameter is smaller than that of coarse slag. The inorganic components of coarse slag and fine slag are mainly SiO₂, Al₂O₃, CaO and Fe₂O₃, in which the acid oxides account for about 35%−80% of the total components, the basic oxides account for about 20%−65% of the total components, and the coarse slag has more basic oxides, and the fine slag has more acidic oxides. The crystalline minerals and amorphous substances in coarse slag and fine slag are not simple coal gasification residues, but formed after a series of complex physical and chemical reactions in the gasification process. Both of the main crystalline minerals of coarse slag and fine slag are quartz, calcite and mullite, but the types and contents of other crystalline minerals are very different, which is mainly affected by factors such as coal type and gasification temperature. The carbon residue content of coarse slag is mostly 3%−20%, while that of fine slag is mostly 20%−40%. The carbon residue content of coarse slag is lower than that of fine slag, but the reactivity of coarse slag is higher than that of fine slag. The distribution of carbon residue content in coarse slag or fine slag with different particle sizes is uneven. The medium coarse slag with about 0.250 mm has a higher carbon residue content, while the highest carbon residue content in fine slag corresponds to the particle size of about 0.125−0.250 mm or large particle size, and the carbon residue content in fine slag increases with the increase of particle size. The residual carbon in the gasification slag has a porous structure with more large pores and good connectivity. The residual carbon in the gasification coarse slag is mostly amorphous, while the residual carbon in the gasification fine slag has obvious aromatic structure and more aromatic C—C or C—H bonds. In terms of heavy metal migration, there are different degrees of heavy metal enrichment in coarse slag and fine slag, and the enrichment factors are closely related to the volatility of heavy metal elements. Compared with coarse slag, fine slag has greater environmental risk, and heavy metals such As Cd, Ni, Cu, Zn, AS, Co, Mo, Se, Pb and Cr in gasification slag have higher environmental risk, which should be paid more attention to. The leaching behavior of heavy metals in gasification slag is dependent on leaching scheme, leaching time and particle size of gasification slag. Under strong acid or/and strong alkali conditions, most heavy metals will have higher leaching concentration, and the leaching concentration of heavy metals increases with the increase of time until it becomes stable. Meanwhile, the leaching concentration of heavy metals in small-size gasification slag is usually higher than that in large-size gasification slag.

The coal-based solid waste filling mining technology is a significant representative technology for achieving green and low-carbon coal mining. Filling the goaf with coal-based solid waste slurry can effectively control the movement of roof strata and prevent derived disasters. The research is conducted on the stability of slurry suspending state, in response to the issue of strong sensitivity of particle suspension in long-distance pipeline transportation of coal-based solid waste filling slurry, which can easily lead to unstable slurry concentration then pipeline blockage or bursts, under the developing trend of intelligent and large-scale coal mining. The aim is to reduce pipe blockage and burst accidents of long-distance filling mining of coal-based solid waste. Firstly, by applying orthogonal experiments, range analysis, variance analysis, and significance tests, the primary and secondary factors influencing the rheological properties of coal-based solid waste filling slurry and their significance are evaluated. The order of primary and secondary influencing factors is determined as slurry concentration > cement addition > fly ash addition, and the order of effect significance is slurry concentration > cement addition > fly ash addition. The mass concentration of coal-based solid waste filling slurry is determined to be 75.2%, with the optimal ratio of cement∶fly ash∶coal gangue∶water as 12∶19.5∶43.7∶24.8. Secondly, a model for solid particle sedimentation based on the solid-liquid two-phase carrier suspension fluid is established for long-distance pipeline transportation of coal-based solid waste filling slurry. The density of the solid-liquid two-phase carrier suspension fluid is determined to be 1626.82 kg/m³, through the calculation of critical non-settling particle diameter mechanical model, granularity screening experiment, and rheological characteristic parameter test. The coal-based solid waste filling slurry is identified as an unstable slurry, lacking long-term suspension stability for long-distance pipeline transportation. Finally, the critical suspension yield stress value and critical suspension plastic viscosity value required for the coal-based solid waste filling slurry to transition into a suspension-stable slurry are determined to be 172.87 Pa and 2.39 Pa·s. A method is proposed to establish a stable slurry system for long-distance pipeline transportation of coal-based solid waste filling slurry by adding a suspending agent, with the aim of solving the practical engineering problems of long-distance filling mining of coal-based solid waste in coal mines.