Accurate identification of the grouting position in overburden isolated grouting and the reasonable determination of the protective layer thickness are of significant importance for preventing surface subsidence and protecting surface buildings and infrastructures. In order to realize the accurate grouting and reduce subsidence in multi-bed separation grouting. By improving the method for determining separation positions based on the composite beam theory, the filling positions for multi-bed separation grouting were determined. Using the theory of elastic thin plates, a structural mechanical model of the protective layer for overburden isolated grouting was established. An equivalent model for the grouting filling space of overburden isolated separation was constructed, and the minimum and maximum total volumes of the grouting filling space were determined according to the breaking law of the upper and lower rock layers in the grouting space. The effect of multi-bed separation grouting in “three soft” thick coal seams was evaluated through the use of probability integrals, combined with actual surface monitoring results. Additionally, 22151 working face of Peigou Coal Mine in Zhengzhou mining area was analyzed as case study. The results show that field borehole observation validated the accuracy of the bed separation positions. Multi-bed separation grouting was successfully implemented between the bottom of medium grained sandstone 152.6 m away from the coal seam and the bottom of fine-grained sandstone 168.6 m away from the coal seam. When the mining thickness of the working face is 7.1 m, a protective layer thickness of 25.2 m ensured the safe implementation of overburden isolated grouting. Theoretical grouting volume calculations range from approximately 9.5 × 10⁴ to 14.7 × 10⁴ tons while the actual engineering grouting volume for multi-bed separation grouting is approximately 13.3 × 10⁴ tons, with a grouting-to-mining ratio of about 0.41. The maximum surface subsidence is 649.8 mm, and the maximum horizontal deformation value of the surface at most residential buildings is within the Ⅰ degree of damage. The surface subsidence control effect of mining with multi-bed separation grouting was remarkable, and the subsidence reduction rate was about 77.91%.

During the mining of the adjacent working face, the strong mine pressure caused by the hard thick roof has obvious zoning characteristics due to the mining of the previous deeply buried large mining height working face. The mechanism of the strong mine pressure at the adjacent working face is the basis for realizing safe and efficient mining. This paper takes the deeply buried large mining height working face at Huangling No.2 Mine as the background, and adopts the methods of geological investigation, theoretical analysis and numerical calculation to analyze the characteristics of the strong mine pressure, the breaking characteristics of the hard thick roof, and the evolution relationship between energy and stress, so as to reveal the occurrence mechanism of strong mine pressure at the adjacent working face of the deep-buried large mining height, and specifies the prevention and control direction of the strong mine pressure. The results show that the number and energy of microseismic events generated by the hard thick roof accounted for 52.12% and 69.4% of the total number of events and total energy, respectively. Zoning display feature was formed in which the value of the energy release on the proximal side was large, and the interior of the roof plate on the solid side was basically intact along the tendency of the working face. The hard thick roof plate constraint boundary conditions of “two solid support, one simple and one free” breaks at the working face tends to be about 108m from the coal pillar of the adjacent section of the breakage. The energy and stress of the hard thick roof under the mining disturbance of 21422 working face form a reduced distribution characteristics from the airside to the solid side. The zonality of the stress and energy distribution is consistently related to the asymmetric breaking of the roof, as derived from the concentration difference coefficient. The hard thick roof has a high degree of release in the 110m range on the air side, and is in a energy storage stage in the range of about 70m from the coal pillar on the boundary of the solid side. According to the results, the strategy of weakening the stress concentration of the roof plate on the airside and blocking the transmission of stress on the solid side has been formed, which provides a reference basis for the effective prevention and control of the strong mine pressure in the deeply buried large mining height working face.

Natural rock mass is not a uniform continuum due to discontinuous interfaces such as faults and bedding planes between rock layers, and the propagation law of acoustic emission signals will inevitably change when they pass through faults and bedding planes. Therefore, studying the propagation law of acoustic emission signals in faults has become one of the key topics in rock mechanics. Based on the Huygens principle, the wave front equation under the condition of heterogeneous media containing faults was derived, and 45°, 60°, 75° and other types of fault specimens were made through laboratory similar simulation model tests. The acoustic emission signals across faults were monitored and recorded by combining ultrasonic tachymeter and DS5-16B full-information acoustic emission signal analyzer. Nonlinear fitting and numerical calculation of Matlab software are used to study the influence of faults and the number of bedding planes of different inclination angles on the propagation speed and signal characteristics of acoustic emission signals. The results show that the propagation speed of acoustic emission signals increases gradually with the increase of fault dip angle, and the propagation speed is positively correlated with fault dip angle. The larger the fault dip angle is, the faster the signal propagation speed will be, and the propagation speed will attenuate after the signal passes through the fault, and the larger the fault dip angle is, the smaller the proportion of velocity attenuation will be. The propagation velocity is attenuated by the bedding plane, and the single bedding plane has little influence on the velocity, while the two bedding planes have great influence on the velocity. The fault will make the maximum value of the signal decrease, the main frequency decrease, and the frequency interval move to the low frequency direction. The larger the fault inclination, the larger the maximum value, the main frequency and the frequency interval. One layer has little influence on the signal, which is basically the same as the time-frequency characteristics of the signal without stratification, while the two layers have a greater influence on the signal, which will greatly reduce the maximum value, main frequency and frequency interval of the signal. The existence of fault will cause the instantaneous energy of acoustic emission signal to decrease greatly, and the smaller the inclination angle, the more serious the attenuation is. The research results can provide theoretical basis for the establishment of wave velocity model under the ray theory.

To investigate the effects of roof conditions, sidewall conditions, and support conditions on roof deflection variations in ultra-long working faces, and to apply beam model theory to hydraulic support design, this study is based on the ultra-long working face 132202 of Xiaobaodang No.2 Mine. A two-dimensional continuous beam model supported on elastic supports was established, reflecting the relationships among hydraulic supports, sidewalls, and roofs in actual production. Using the displacement method, beam model elements were encoded, and the element stiffness matrix for each beam segment was calculated. By employing the element assembly method, the global stiffness matrix and equivalent nodal loads of the ultra-long beam model with elastic supports were computed. Through the matrix displacement method, the global deflection distribution of the beam and the end forces of beam elements were derived, enabling the calculation of support reaction forces. Parametric analyses were conducted on variables such as support width, number of supports, equivalent stiffness, working face length, sidewall stiffness, roof elastic modulus, moment of inertia, and loads induced by adjacent strata. Their impacts on the global deflection distribution were examined. A cubic polynomial was used to precisely fit the initial-to-peak segment of the deflection curve on one side. Validation was performed using results from 3DEC numerical simulations incorporating pile structural elements for support, alongside field monitoring data from electro-hydraulic control systems. The validation confirmed consistency among theoretical calculations, numerical simulations, and field data, demonstrating that the 2D beam model reasonably explains the tri-peak loading characteristics observed in ultra-long working faces. Additionally, the deflection curve from the edge of the working face to the adjacent peak value aligns with a cubic polynomial distribution. This study deepens the application of beam models in mining and provides guidance for hydraulic support design in ultra-long working faces.

The coal mine environment is easy to cause corrosion damage to the anchorage material in long-term use because its complexity, which threatens the safety of roadway support. To characterize the corrosion characteristics of coal mine environment, the corrosion factors and corrosion tendency of coal mine environment were studied. The environmental characteristics and main corrosion factors of coal mine were analyzed based on field investigation. The multi-parameter coupling evaluation method of coal mine environmental corrosion tendency is designed by analytic hierarchy process-entropy combination weighting and membership degree analysis. The environmental corrosion tendency grade of underground roadway in Gaojiabao Coal Mine was determined by industrial application test. The results indicate that the coal mine environment can be divided into three categories: water environment, atmospheric environment and surrounding rock environment, including 9 main environmental corrosion factors. By index weighting analysis, the weight of water environment was determined as 0.64829562, the weight of atmospheric environment was determined as 0.22967309, and the weight of surrounding rock environment was determined as 0.12203129. Nine influence indexes of main environmental corrosion factors in coal mines were put forward by membership analysis, such as pH influence index, Cl⁻ influence index and ${\mathrm{SO}}_4^{2-}$ influence index. The membership function of each main environmental corrosion factor was constructed. Combined with index weighting and membership analysis, the multi-parameter coupling evaluation method of coal mine environmental corrosion tendency was designed, the environmental corrosion tendency of coal mine is comprehensively judged by environmental detection, index quantification, evaluation index determination and engineering criterion comparison. The environmental corrosion tendency of the north wing return air roadway, the west wing concentrated auxiliary transportation roadway, the west return air roadway, the 3403 transportation roadway and the 3407 transportation roadway in Gaojiabao Coal Mine were determined as grade V, grade III, grade II, grade III and grade IV by applying the multi-parameter coupling evaluation method of coal mine environmental corrosion tendency. Comparing the results of roadway environmental evaluation and industrual corrosion test, it can be seen that the evaluation method can accurately evaluate the corrosion characteristics of coal mines environmental. This study can effectively evaluate the mine environment and provide guidance for the anti-corrosion support of coal mine roadway.

Under the excavation effects of deep coal and rock, obvious creep damage is produced under the action of high stress for a long time and the sudden destabilization of the creep-damaged coal body induces rock burst disaster under the mining action in working surface, which poses a serious threat to the safety of deep coal pillar mining. Through long-term creep-cyclic loading-unloading confining pressure test, the evolution law of deterioration characteristics of coal rock under the superposition of initial creep damage and cyclic load is studied. The strength characteristics and cumulative damage characteristics of coal in different stress intervals under cyclic load are analyzed. The coupling mechanism of cyclic load times on creep damage effect of coal is explored, and the energy conversion mechanism in the process of deformation and failure of coal under creep damage effect is revealed. The research results show that: coal samples are less affected by creep pre-damage effect within the low stress range of cyclic loading, with the increase of the stress level range, the larger the long-term damage of coal samples, the more obvious the nonlinear deterioration, and the more severe the degree of rupture after sudden destabilization; When the creep stress is in the elastic stage, the time between strengthening and deterioration is generally less than 16 h, and in the high stress range, the deterioration effect is significantly; At the later stage of cyclic loading and unloading, the loading and unloading deformation modulus fluctuates sharply with the number of cycles, and the irreversible deformation increases steadily, indicating that the coal samples is at the “critical point” of sudden destabilization. Based on the theory of damage mechanics, a damage evolution model of creep and cyclic loading is established. Combined with the test results, it is found that the internal deterioration of coal samples after long-term creep is high, the stored energy and stress release are less, and the sudden instability is less evident; The energy storage capacity of coal samples also has a “critical point” in terms of creep damage duration. Due to the longer creep damage duration, the releasable elastic energy stored in the specimen is reduced, which retards the release of elastic energy to a certain extent, reducing the scope of the occurrence of dynamic disaster damage. The research results will be of great engineering significance to reduce the long-term creep induced disaster of residual coal, improve the recovery efficiency of residual coal in goaf, and promote the construction of ecological civilization in mining area.

Abstract: In order to study the macro and micro damage characteristics of fractured rock mass under cyclic load, cyclic loading and unloading tests with variable upper limit were carried out to analyze the mechanical properties and macro failure characteristics of the samples. Based on the results of XRD analysis, a UDEC numerical model of fractured sandstone is constructed to analyze the microscopic damage law of the model and reveal its mechanism. The results show that: loading and unloading can strengthen the deformation resistance of the samples with small cracks, and on the contrary increase the failure velocity of the samples with large cracks; With the increase of the prefabricated fracture length, the macroscopic failure characteristics of the sample changed from tensile shear failure to shear failure, and the irreversibility of the damage gradually increased. The damage evolution process of the sample under loading and unloading can be divided into four stages: "elasticity-strengthening-fatigue-failure". No damage occurs in the elastic stage model, while an effective resistance structure is formed in the strengthening stage model. Cumulative damage increases significantly in the fatigue stage, and tensile cracks expand during unloading. During the loading stage, cracks expand and connect with each other, the contact point pole of the block between adjacent shear microcracks is destroyed, and the connection between the blocks fails. During the unloading process, the stress drop model is relaxed, and the failed connection between the blocks opens up at this time, leading to the initiation of tension cracks between the shear microcracks, and accelerating the instability and failure of the model. The research results reveal the macro and micro damage characteristics of fractured sandstone under variable upper limit cyclic loading and unloading, and provide a reference for scientific evaluation of engineering stability.

The development and expansion of fractures in coal rocks are the primary causes of mining disasters, such as gas outbursts, rock bursts, and instability of coal pillars. Investigating the fracture evolution process and damage evolution law in coal rocks is crucial for predicting dynamic hazards in coal mines. Fracture inclination angle and porosity are significant factors influencing the propagation behavior of fractures in coal rocks. In this study, we conducted uniaxial compression tests, CT electron scanning, and nuclear magnetic resonance experiments to analyze the mechanical properties and pore-fracture structure of coal rocks. We characterized the fractal dimension of internal fractures in coal rocks by analyzing their distribution characteristics along with fracture inclination angles. Additionally, numerical simulations were employed to analyze the propagation and damage evolution laws of fractures with different inclination angles in coal rocks. The research findings demonstrate that: ① The degree and distribution characteristics of fracture inclination angles significantly impact the complexity and fracturing behavior of coal rock; when a single dominant fracture exists, steeply inclined coal rock exhibits higher compressive strength compared to gently inclined or nearly vertical ones; compressive strength is higher for single-fractured coals than complex-fractured ones. ② Porosity is the primary factor influencing the compressive strength of coal rock of the same type. As porosity increases, the compressive strength of coal rock decreases and exhibits a direct correlation with the presence of large pores. ③ Single-fracture coal and rock exhibit a lower level of complexity compared to complex fractured ones. This results in reduced stress concentration during compression, leading to less initial damage. However, during accelerated crack propagation, there is a rapid release of stored strain energy, resulting in higher levels of shear failure. ④ By establishing a variable for coal and rock damage based on the number of developed fractures, it can be observed that the evolution process follows an exponential growth pattern. This process can be divided into stages including approximate intactness, initial damage, stable development of damage, accelerated development of damage, and residual damage. Such observations provide theoretical support for predicting mining disasters.

The stress environment of deep coal-rock mass is complex, to further reveal the mechanism of rock burst induced by dynamic load. Based on the Split Hopkinson Pressure Bar test system, the dynamic tests under different stress states (non-axial and confining loads, one-dimensional coupled static-dynamic loads, three-dimensional coupled static-dynamic loads) and strain rates (49.3～137.9 s⁻¹) were carried by using the rock-coal-rock structure samples. The characteristics of strength, deformation, failure and energy evolution of rock-coal-rock structure samples were studied. The results show that there were two types of stress rebound and strain softening after the peak of the stress-strain curve of the composite specimen under non-axial and confining loads and under one-dimensional coupled static-dynamic loads, and there was stress rebound phenomena under three-dimensional coupled static-dynamic loads. Under three stress states, the peak strength of the samples increases roughly with the increase of strain rate, showing an obvious rate correlation. Under three stress states, the proportion of reflected energy to incident energy was the highest, and the proportion of transmitted energy to incident energy was the lowest. The proportion of reflection energy to incident energy under three-dimensional coupled static-dynamic loads was lower than the other two stress states. When the strain rate is lower than 123 s⁻¹, the energy utilization rate and dissipated energy increase gradually with the increase of the strain rate, and the dissipated energy density increases with the increase of the incident energy. The failure modes of the samples show an obvious rate correlation, and the size of coal and sandstone fragments decreased gradually with the increase of strain rate. Under non- axial or confining loads and one-dimensional coupled static-dynamic loads, the samples firstly failed at the coal-rock interface. Many small size fragments of coal and sandstone are secondary cracks caused by the effect of loads. The fracture of coal and rock mass is not obvious under three-dimensional coupled static-dynamic loads.

In order to study the time-dependent characteristics of filling paste under the action of mine water, chloride salt solutions with mass fraction of 0%, 5%, 10%, and 15% were prepared, and chloride salt dry wet cycle tests with erosion cycles of 4, 8, 12, and 16 times were carried out. The macroscopic and microscopic characteristics of the filling paste were analyzed, the damage curve of the filling paste was obtained based on the constructed compaction-elastoplastic constitutive model, and the stress evolution mechanism of the filling paste under the action of chloride salt was discussed. The results indicate that the mass of filling paste shows a sharp increase, a slow increase, and a slow decrease trend with the increase of chloride erosion cycles. High-concentration chloride salt solution accelerate the quality change of filling paste. As the cycles of chloride erosion increase, the filling paste exhibits a macro-mechanical characterized by high stress-low strain, low stress-high strain, and low stress-low strain. The compaction degree exhibits a dynamic evolution characterized by an initial sharp decrease followed by a stable variation, the plasticity factor demonstrates a developmental trend of initial stability followed by a sharp change. The chloride salt promotes the stable development of the damage process of filling paste and inhibits the surge of damage in the later stage of plasticity. The development curvature of the damage curve after 16 dry-wet cycles in 5 % and 10 % chloride salt solution is relatively small, and the development curvature of the damage curve after 12 dry-wet cycles in 15 % chloride salt solution is relatively small. Chemical corrosion is a significant factor leading to the deterioration of the binding properties of filling paste. The salt corrosion products from chemical corrosion partly originate from the chemical combination of chloride and unreacted tricalcium aluminate(C₃A), and another portion arises from the chemical bonding of chloride with the hydration product ettringite(AFt). The coordination deformation between salt corrosion products and internal structure is a key factor causing the alienation of the bearing performance of filling paste. The crystalline expansion force of salt corrosion products resists internal stresses of filling paste, resulting in a reduction in compaction performance and crack propagation ability. This study can provide a theoretical basis for the analysis of the time-dependent stability of filling paste in mine water, and this study is of great significance for maintaining the long-term stability of filling paste.

Due to multiple factors, open pit mining will form a large amount of trapped coal on the side wall. With the end of internal discharge and reclamation, these trapped resources will be permanently lost, resulting in great waste. In order to reduce resource waste, shaft mining is widely used, but due to the typical deformation, failure and instability of slope, it brings great safety risks to the mining on the side wall and seriously restricts the recovery of resources. Therefore, it is an important topic for geotechnical engineering and mining engineering to clarify the sliding mechanism and control mechanism of mining slope under mining disturbance and guarantee the stability of slope. In view of this, this paper comprehensively uses physical and numerical simulation, mechanical analysis and other means to study the overburden migration law, the deformation and failure characteristics of mining slope, the instability sliding mechanism and its control mechanism. The following innovative results have been achieved: The evolution law and formation mechanism of the failure form in the “transverse three zones” of the mining slope are revealed, and the relationship between the failure and the coupling failure and instability slip of the mining slope is clarified, and the phenomenon of “circular arc” sliding surface, shear slip zone and local instability slip of the slope is easily caused by mining in the lower coal seam. The co-existent failure characteristics of step collapse and tensile and shear crack of mining slope are obtained. A mechanical model of “masonry beam” of mining slope was established, and the dynamic evolution of “masonry beam” and the relationship between the rotation mechanism of key rock B and the instability and sliding of mining slope were revealed. It is found that the main reasons for the sliding of mining slope are the repeated movement of the “masonry beam” structure, the continuous forward movement of the “transverse three zones” structure, the horizontal thrust formed by the key block B's instability, and the sliding mechanism of “mining cave-bedding creep−slope instability”. The control method of mining slope instability slip for controlling the rotation of key rock block B of “masonry beam” is proposed and analyzed. It is concluded that the backfill pressure foot control method can effectively change the fracture degree of hinged rock block and inhibit the rotation of key rock block B on the basis of curbing horizontal thrust transmission, and achieve the effective control of mining slope instability slip. Based on this, the engineering application and verification are carried out, and the stability control system of mining slope is formed, which provides guarantee for the safe and efficient recovery of side coal.

In view of the problems of high risk in rock cross-cut coal uncovering of deep outburst coal seam, and borehole blowout and low efficiency during underground drilling, based on numerical simulation and field experiments, the mechanism of reducing outburst by cavity completion in surface boreholes for assisting rock cross-cut coal uncovering in high outburst coal seam was studied, and the key process parameters were optimized. Firstly, the multi-stage cavity formation technology and supporting system for surface cluster wells were developed. Then, the evolution of physics of stress-relief coal seam was revealed, and the optimization method of key process parameters for cavity creation was proposed. Finally, the technology of cavity completion in surface boreholes for assisting rock cross-cut coal uncovering was successfully implemented in the field, and it proves that this technique is effective. The research results show that: A multi-field collaborative outburst prevention method of “well network pressure relief-borehole group energy dissipation-curtain solidification” was proposed. The mechanism of outburst prevention by cavity completion in surface cluster well was revealed. A multi-stage caving technology of “borehole mechanical reaming-hydraulic jetting-water drainage pressure relief” was developed, and an integrated system of “cavity creation-slag discharge-water slag separation” was also developed. The range of plastic zone around the cavity is linearly related to the volume of the cavity, with a ratio of 81.90. The gas pressure in the low permeability area near the cavity decreases first and then increases with the increase of cavity diameter, and the corresponding optimal cavity diameter is 2.0 m. When the roadway is arranged along the direction perpendicular to the maximum principal stress, the outburst risk of coal on both sides of the roadway are easier to be eliminated after cavity completion. The technology of cavity completion in surface boreholes for assisting rock cross-cut coal uncovering was successfully implemented in the field, and the results show that the permeability coefficient of the coal seam has increased by about 10 times. The number of boreholes for in the test site has decreased by 24%, and the drilling length has decreased by 21%. The technical method described here can be further expanded to form the “coal-gas co-extraction method with borehole”, which is expected to achieve synergistic and efficient co-extraction of coal and gas in difficult to extract coal seams.

With the gradual deepening of coal mining, the issue of high ground stress will be faced. In response to the unclear understanding of the internal damage and fracture evolution of deep wells with high gas and low permeability coal seams under the influence of deep ground stress, a theoretical model of convergence energy blasting jet formation was established. The critical collapse velocity of the convergence jet formation was studied, verifying the feasibility of jet formation, and analyzing the mechanism of convergence energy blasting load erosion on coal seams. Subsequently, through laboratory experiments, comparative simulations of convergence energy blasting and conventional blasting at different burial depths were conducted, analyzing the effects of convergence energy blasting and conventional blasting on coal seam crack propagation from three aspects: macroscopic cracks, strain conditions, and ultrasonic data. Then, numerical simulation software was used to study the distribution characteristics of coal damage under convergence energy blasting loads at different burial depths. Finally, the effects of normal blasting and convergence energy blasting under different buried depths on gas concentration and gas extraction purity are studied through field tests. The results of blasting similarity simulation experiments show that deep ground stress can inhibit the fracturing range of blasting loads on coal bodies, and conventional blasting is more affected by ground stress than convergence energy blasting. Relative to conventional blasting, the fracturing range of convergence energy explosives in deep coal bodies is wider during permeability enhancement blasting. The maximum pressure strain peak at measurement point 1# in the convergence direction is 1.35 times that of conventional blasting, while at measurement point 3# in the non-convergence direction, it is 86% of conventional blasting. This also results in cracks preferentially propagating in the convergence direction, with less expansion in the non-convergence direction, leading to better directional blasting effects. Numerical simulation results indicate that with increasing burial depth, the range of coal penetration by convergence energy blasting decreases. A relationship formula between different burial depths and the fracturing range of convergence energy blasting was fitted, which can serve as a basis for determining the fracturing range of deep mine convergence energy blasting to some extent. Field test results show that deep hole convergence energy blasting can increase gas extraction to a greater extent than common deep hole blasting. The buried depth has a certain influence on the crack range of deep hole convergence energy blasting, so the influence of ground stress should be considered when using shaped charge blasting in deep mine. The research findings have certain reference significance for the blasting permeability enhancement of deep high gas and low permeability coal seams.

The freezing coring technology can not only perform the fixed-point sampling in coal seams, but also effectively reduce gas leakage during the sampling process. It has broad prospects for application in the precise measurement of gas content in coal seams. To accurately evaluate the gas loss of coal sample during the freezing coring process, the simulated tests of gas desorption during coring were carried out under the different external heats and the equilibrium pressures (1−4 MPa) by using a simulation platform for gas adsorption/desorption on gas bearing coal. The gas desorption characteristics in low-temperature conditions were also studied. Based on the graphical method, the three diffusion models were adopted to analyze the gas desorption curves under low-temperature conditions, and the fitting performance of the gas loss estimation models were evaluated. The results show that the tube frictional heat generated during the conventional coring process greatly increases the gas loss; and the gas desorption amount gradually increases with the temperature rise in the core tube wall. When the friction heat of the tube wall is 60, 70, 80, and 90 ℃, the desorption amounts within 30 min are 6.587, 7.082, 7.460, and 7.981 cm³/g, respectively; and the increments in desorption amounts compared to that of 30 ℃ are 13.71%, 22.25%, 28.78%, and 37.77%, respectively. During the freezing coring, there is a back-flow phenomenon in the desorption, which is caused by the coal sample cooling leading to the pressure inside the sample tank lower than atmospheric pressure. Through the correction tests of the pure back-flow of coal without gas, the gas loss during the freezing coring gradually decreases as the external heat of the tube wall. When the external heat of the tube wall is 60, 70, 80 and 90 ℃, the desorption amounts within 30 min are 3.578, 3.842, 4.215, and 4.76 cm³/g, respectively; and the desorption inhibition rates reach 40%−46%, compared with the conventional coring. At low temperatures, the desorption amount gradually increases with the rise of adsorption pressure, but the growth rate gradually decreases. The diffusion coefficient during the freezing coring is much reduced compared to the conventional coring, and it shows a linear decrease with cooling. The fitting accuracy of the gas desorption curve at low temperatures via the logistic growth model is significantly better than that of the $\sqrt{t}$ model and the exponential model, with a loss estimation error of less than 0.5%. Therefore, the logistic growth model can meet the needs of gas loss estimation during the freezing coring in coal seams.

The non-sealing pressure measurement technology based on “state recovery” is a new experimental pressure measurement technology that compensates the gas loss, restores the coal core volume, and artificially simulates the coal reservoir temperature to restore the coal seam gas pressure under the original coal body occurrence state by transferring the gas pressure test coal sample from the original coal body occurrence state to the stripped coal sample and put it into the experimental coal sample tank. It is a feasible new technology to obtain the residual gas pressure of pre-pumped coal seam accurately and quickly, which is not limited by the location of downhole pressure measurement and the quality of hole sealing.The accurate recovery of the core volume is the key to the accurate determination of residual gas pressure in the pre-pumped coal seam by this method. In order to explore the appropriate coal core volume restoration method, relying on the no-seal coal seam gas pressure tester independently developed by the team, the coal core volume restoration experiment was carried out by loading axial compression coal core volume restoration, contact liquid injection (water/oil) coal core volume restoration, and the combination of pressure and liquid injection coal core volume restoration. The results showed that: When the loading axial pressure is set as 5, 10, 15……65 MPa in order to restore the coal core volume, due to the mutual support and resistance between the coal sample particles, the newly formed cracks caused by the sampling of broken coal body cannot be completely "healed", and it is difficult to restore the coal core volume by simply pressuring it in both theory and technology. Set the initial adsorption equilibrium pressure P₀ to 0.31, 0.53, 0.74, 1.00 and 1.50 MPa for the coal core volume recovery test of contact water injection, contact silicone oil injection and compression/liquid injection combination. The experimental error of the coal core volume restoration method is 40.67%～116.13%. The experimental error of the coal core volume restoration method with contact silicone oil injection is 29.33%～87.10%, and the experimental error of the coal core volume restoration method with pressure/liquid injection is 2.67%～9.68%, and the error decreases with the increase of adsorption equilibrium pressure. Comparing the experimental results of three kinds of coal core volume restoration methods, it is finally determined that the combination of pressure in the upper cavity and injection of silicone oil in the lower cavity restores the coal core volume in situ. It provides a reliable way for the realization of non-sealed hole pressure measurement technology based on “state recovery”, provides a new method for coal mine to accurately and quickly obtain the residual gas pressure value of pre-pumped coal seam, and provides technical support for coal seam gas extraction effect inspection and extraction standard evaluation.

In order to eeproduced as much as possible the disaster processes such as air volume fluctuations, airflow reversal, abnormal airflow temperature, the spread of toxic and harmful smoke, and the decrease in oxygen volume fraction in the ventilation network during the period of fire occurrence, the airflow pressure balance equation of the ventilation network circuit based on the conservation of airflow mass flow rate, the control equations for heat exchange and propagation in ventilation networks, the control equations for the diffusion and dispersion of toxic and harmful gases in ventilation networks, and the control equation of the consumption and transport of oxygen component in ventilation network were constructed. Based on the theory of multi physics field coupling, the above equations were combined using key ventilation parameters such as airflow resistance, fire pressure, and air flow density to construct the multi field coupling mathematical model for an unstable ventilation network during the period of external fire, which included “air volume and airflow direction, airflow temperature, toxic and harmful gases volume fraction, and oxygen volume fraction”. Based on the Newton method of loop air volume, the variation of air volume and airflow direction in the ventilation network was solved. The finite difference method of upwind discrete format was used to solve the heat and mass transfer process of airflow in the ventilation network. The indirect coupled multi physical field solution method based on temporal logic was adopted to construct the solution process. The accuracy and applicability of the model were preliminarily verified by carrying out external fire experiments in the test roadway of the angular structure, and on this basis, the dynamic change process of air volume change, wind flow reversal, abnormal wind temperature, spread of toxic and harmful gases, and oxygen volume fraction fluctuation in the complex ventilation network under the location of the fire source of the branch in the air inlet area and the branch branch in the air consumption area were simulated respectively.The simulation results showed that under the condition that the ignition source was located in the air inlet well, the number of branches with significant changes in air volume reached 61% of the total number of branches in the ventilation network, the number of branches with significant changes in wind flow reached 11% of the total number of branches in the ventilation network, the number of branches with significant changes in wind temperature reached 76% of the total number of branches in the ventilation network, the number of branches with CO intrusion reached 84% of the total number of branches in the ventilation network, and the number of branches with significant changes in oxygen volume fraction reached 41% of the total number of branches in the ventilation network. The impact scope and intensity of disasters caused by fire in the air inlet shaft branch were significantly greater than those in other branches.

The composite disaster of gas and coal spontaneous combustion in the goaf is increasingly becoming the main disaster pattern that restricts the safe production of mines. To study the disaster mechanism of the composite disaster of gas and coal spontaneous combustion, the experiments of low-temperature oxidation microstructure and thermal analysis of coal under different CH₄ adsorption pressures were carried out in the paper by low-temperature liquid nitrogen adsorption, scanning electron microscopy and TG-FTIR testing methods, and a pore fractal calculation model was set up based on the fractal theory to investigate the influence of adsorbed CH₄ on the pore structure and thermal stability of coal. The results showed that the adsorbed CH₄ inhibited the thermal damage destruction of micropores by oxidation, thus a large number of micropores existed in gas-bearing coal and the pore morphology changed compared with the raw coal which was mainly medium- and large-porous; with the increase of adsorption pressure, the mass loss of coal samples in the stage of water loss and desorption was 4%, 2.9%, 3.2% and 3.2%, respectively, and CH₄ occupied the adsorption sites of the coal oxygen reaction leading to the reduction of compounds involved in the oxidation reaction, and the active substances involved in the oxidation reaction decreased in the stage of oxidative weight gain, and the oxidative weight gain of the coal samples was 0.67%, 0.41% and 0.31%, respectively, and the spontaneous combustion of the gas-bearing coal in all stages of the characteristic temperature points lagged with the increase of adsorption pressure. A fractal model based on pore size distribution and CH₄ adsorption/desorption process was established. The morphological inhomogeneity of the coal surface increased during the pre-oxidation period of low CH₄ pressure, the gas adsorption capacity of the coal surface was elevated, the possibility of oxidation reaction increased, the risk of coal spontaneous combustion increased, and the ability of low gas to inhibit the spontaneous combustion of coal was weaker, and was correlated with the specific surface area of the primary pores. The research results investigated the influence of different residual gas contents on the spontaneous combustion characteristics of coal, further verified the accuracy of the continuous physical simulation platform for the whole process of adsorption-desorption-oxidation of gas-bearing coal competition under different adsorption pressures and temperatures, and provided basic support for the prevention and control of gas-fire coupling disaster in the composite disaster environment of the goaf.

Rapid prediction of flow loss in goaf is always the key problem of accurate fire prevention in coal mines. Based on the porous media flow in goaf, a physical model of porous media in goaf assembled by spheres is proposed. Combined with the geometric characteristics of the model, the minimum volume element of the flow is found, and four stages of the flow are divided. Based on Navier−Stokes equation, the relationship between apparent velocity, diameter, dynamic viscosity, fluid density and pressure loss is established, and the pressure loss calculation formula of four stages is obtained. Further, the pressure loss model per unit length J_T/L is established by using the time-specific weight method, which accords with Forchheimer equation form. The model is used to predict the pressure loss, and the simulation results are compared with the predicted values of Ergun equation. The results show that the boundary effect on the numerical simulation results is greater than that of the inlet and outlet effect. The optimal model is 16×7 multi-ball assembly unit grafting combination. For the porous media flow in the grafting combination, the relative error between Ergun predicted value and simulation result is less than 8% in laminar flow or near laminar flow. The reliability of the numerical simulation results was verified through dimensionless analysis using experimental data from porous media with a porosity of 0.375. However, the maximum error of J_T/L predicted value of the unit length pressure loss model is 69.9%, which is actually caused by the simplification of N−S equation to one dimension. The correction coefficient K was defined to represent the influence of three-dimensional flow. When the Reynolds number (R_e) is less than 10, K=2.84, and when the R_e is from 10 to 3000, K=6.56R_e^−0.334. Finally, the pseudo-three-dimensional J_T'/L model is obtained. When the R_e is below 10, the error of the predicted value is less than 6%, and when the R_e is from 10 to 3000, the error is less than 10%. It is concluded that the pseudo three-dimensional porous media flow pressure loss calculation model based on the sphere assembly model can quickly predict the pressure drop per unit length of the R_e less than 3000, which is helpful to scientifically guide the engineering practice of accurate fire prevention in the goaf.

Horizontal well fracturing stimulation is currently the most core technical means to achieve efficient development of unconventional coalbed methane resources. Although good on-site results have been achieved, there are still some problems such as unclear the goal of fracturing stimulation, unclear the adaptability of the main horizontal well fracturing technology and lack of evaluation methods for fracturing effects. In order to further improve the production and development benefits of fractured horizontal wells, this article based on the viewpoint of “unblock and channeling”that has achieved significant results in the southern part of the Qinshui Basin and has been verified by practice, combined its requirements for fracturing on high-rank coal reservoir and the characteristics of the original natural pores and fractures in the high-rank coal reservoirs in the southern part of the Qinshui Basin, referred meanwhile to the evaluation experience of previous fracturing, selected seven key indicators: Types and quantity of fracturing operation curves for each section of the horizontal well, average operating pressure, instantaneous shut-in pressure, accident type and occurrence rate, Pressure at the moment of initial gas production, Current bottomhole flowing pressure and daily gas production (Production time is more than one year), and comprehensive average stable gas production, proposed an evaluation method for the fracturing effect. Furthermore, combined with on-site comparative experiments, a study on the adaptability of the main horizontal well fracturing technology will be conducted to determine the upper limit and shortcomings of the main fracturing technology’s stimulation capacity. The results show: With the increase of burial depth and geo-stress, the difficulty of coal reservoir stimulation has shifted from easy to filter out, prone to excessive opening of natural fractures near the wellbore, making it difficult to create dominant main fracturing fractures, to difficulty in gradually opening natural fractures around the dominant main fracturing fractures, insufficient opening of secondary fractures, and difficulty in transporting fracturing sand, making it easy for desanding and sand-block.Due to the excessive pressure loss along the wellbore, the bottom sealing drag fracturing technology is unable to meet the needs of “unblocking and channeling”fracturing stimulation at depths of more than 1 000 meters.The continuous tubing fracturing technology has the problem of insufficient carrying capacity for fracturing sand after at depths exceeding 1 200 meters.The bridge shooting combined fracturing technology can not only meet the needs of the “Unblocking and Channeling” fracturing stimulation of coal reservoirs with a depth of 1 250 meters, but also has the potential to further increase the scale of fracturing stimulation, which can meet the construction and implementation of the“unblocking and channeling”fracturing stimulation of coal reservoirs under deeper conditions. The research provides important basis and support for the rational selection and upgrading of fracturing technologies in the southern part of the Qinshui Basin and other coalbed methane blocks at home and abroad.

A set of organic rich shale with good exploration potential is developed in the Wufeng Formation Longmaxi Formation of the Shaozhai Syncline in northeastern Yunnan; However, its geological characteristics, gas bearing characteristics, and main controlling factors are still unclear. By systematically reviewing previous data and combining with the latest drilling and profile data, the petrological characteristics, organic matter characteristics, reservoir space characteristics, and gas bearing characteristics of organic shale rich in the Wufeng-Longmaxi Formation of the Shaozhai Syncline were analyzed, and the controlling factors of gas bearing properties were discussed. The results show that the Wufeng-Longmaxi Formation in the Shaozhai Syncline is a shallow water shelf deep water shelf deposit, with a thickness of organic rich shale ranging from 35 to 105m; The organic rich shale in the upper section of the Wufeng Formation and the lower section of the Longmaxi Formation has a high TOC content, mostly above 2%, with good organic matter types and good hydrocarbon generation potential; High quality shale has a high content of brittle minerals and good fracturing performance. There are multiple types of reservoir spaces, including shale gas storage pores mainly composed of organic matter pores, intergranular pores, and intra particle pores, as well as mineral fractures and interlayer fractures of clay minerals that communicate with fracturing channels and macroscopic fractures. The gas content of shale in the study area varies greatly. The gas content in the Gaoqiao area is relatively high, but the CH₄ air free base content is low. The gas content in the Lianfeng area is relatively low, but the CH₄ air free base content is high, both of which are above 95%. The study found: the difference in total gas content of shale at the same structural location is controlled by TOC content, pore volume, pore specific surface area, clay mineral content, pyrite content, and carbonate mineral content. The shale gas content at different structural locations is controlled by the sedimentary environment and preservation conditions under structural control. The deep water continental shelf, wide and gentle synclines far from major faults and with fold shielding under structural control have better gas bearing properties, it is a favorable area for shale gas enrichment in the study area.

Based on the frequency dispersion theory, the channel wave can inverse the coal seam thickness distribution in mining face and delineate the thin coal area to a certain extent by using a certain frequency or band-pass filter frequency band. However, due to the complexity and variability of the coal seam thickness in mining face, there is a significant difference between the actual and theoretical dispersion characteristics of channel wave, which reduces the prediction accuracy of coal seam thickness. In order to further improve the effect of channel wave in predicting coal seam thickness, based on the theoretical relationship of group velocity, coal seam thickness and frequency, the sensitive frequency, narrow sensitive frequency band and wide sensitive frequency band of different coal seam thickness are calculated by second-order, third-order and fourth-order derivative, and using the relation between group velocity and frequency, each order derivative and its variation curve. The coal seam thickness and its range can be calculated for different sensitive frequencies. A coal seam thickness prediction method based on multi-frequency joint application of channel wave is proposed by using multiple sensitive frequency weighted reconstruction and sensitive frequency band. For the test mining face, the coal seam thickness is inverted by the group velocity of the 147 Hz sensitive frequency, the narrow sensitive frequency from 123 Hz to 172 Hz, and the wide sensitive frequency band from 103 Hz to 188 Hz corresponding to the disclosed coal seam average thickness in tunnelling. At the same time, considering the effective frequency band of channel wave and the range of mainly disclosed coal seam thickness in tunnelling, the group velocity of each coal seam thickness corresponding to the sensitive frequency is weighted and reconstructed with the proportion of each coal seam thickness as the weight factor, and the coal seam thickness in mining face is predicted by the multi-frequency weighted reconstruction group velocity. The distribution trend of coal seam thickness predicted by single sensitive frequency, narrow sensitive frequency band, wide sensitive frequency band and multi-frequency weighted reconstruction is basically consistent with the actual mining situation, and the thick coal seam area and thin coal seam area are well predicted. The prediction accuracy of the four methods can reach 76.15%, 83.68%, 81.17% and 86.62% respectively for the coal seam thickness with the deviation not more than 20% (0.50 m). The results show that the sensitive frequency and the sensitive frequency band will decrease, and the range of sensitive frequency band will also decrease, while coal seam thickness increases. With the increase of frequency, the predicted optimal coal seam thickness will become thinner, and the thickness range will also become smaller. Single sensitive frequency, narrow sensitive frequency band, wide sensitive frequency band and multi-frequency weighted reconstruction can effectively predict the thickness of coal seam in mining face. However, the prediction accuracy and precision of the sensitive frequency band and multi-frequency weighted reconstruction of multi-frequency joint application are improved compared with the single sensitive frequency. In particular, the multi-frequency weighted reconstruction has the best effect. Therefore, multi-frequency weighted reconstruction can be preferred, narrow sensitive frequency band and wide sensitive frequency band can be followed, and single sensitive frequency is the finally in predicting the coal seam thickness by channel wave.

It is an effective method to reduce the risk of floor water inrush by grouting reconstruction and sealing of water channel on top of Ordovician limestone before deep confined water mining in Huabei type coal mine in China. The basic properties of fly-ash cement materials with different ratios were tested using the orthogonal test method, in accordance with the requirements of deep medium to strong pervious limestone grouting project in Zhuozishan coal field. Subsequently, the influences of three factors, water-solid ratio, solid ratio and temperature, on six indexes of slurry density, viscosity, water extraction rate, setting rate, setting time and mechanical strength were analyzed. The results show that the ratio of water to solid has a relatively prominent effect on 6 indexes of slurry. The solid ratio has great influence on the viscosity, setting time and mechanical strength of the slurry. The effect of temperature on the performance of the slurry is relatively small. The pressure boosting process of ground grouting holes is divided into three stages: low-pressure filling, medium-pressure diffusion, and high-pressure fracture extension and reinforcement. The grouting material ratio varies across different stages of grouting. The comprehensive balance method yielded three material ratio schemes tailored for various grouting stages. The optimal grouting material ratio is applied for the grouting renovation of medium-strong pervious limestone. Engineering practice demonstrates that the slurry effectively diffuses and fills cracks in the formation, essentially intercepting the vertical water channel exposed at the top of the Ordovician limestone. It reduces the risk of water inflow on the working face and provides technical support for regional grouting treatment projects targeting the top of the Ordovician limestone.

Coal gangue sorting is the most basic, effective, and important technical measure to improve coal quality. Improving the accuracy and efficiency of coal gangue sorting is a serious challenge faced by coal gangue sorting. In-depth research and analysis have been conducted on the existing intelligent sorting systems for coal gangue, including “grabbing sorting”, “fork sorting”, and “pneumatic sorting”. The academic thought of “Sorting gangue is sorting images” has been proposed, and the logical framework of the academic ideology of “Sorting gangue is sorting images” has been established. The basic connotation of the academic ideology of “Sorting gangue is sorting images” has been elucidated, mainly including image-based coal gangue recognition, image-based coal gangue sorting feature extraction, image-driven sorting machine dynamic target tracking, and multi-task multi-sorting machine collaboration based on image sequences. Aiming at the problem of image-based coal gangue recognition, a recognition principle and metsorting rates visual images and X-ray images is proposed, which can effectively improve the accuracy of coal gangue recognition; Aiming at the problem of feature extraction in coal gangue image sorting, a plane and depth feature extraction and fusion algorithm based on coal gangue images is proposed. A coal gangue sorting cube is constructed, which can improve the accuracy of coal gangue sorting; An image-based coal gangue matching tracking and path planning method is proposed for dynamic coal gangue tracking, which can improve the accuracy and reliability of sorting; Aiming at the problem of intelligent collaborative sorting of multiple sorters, it is proposed to construct a comprehensive benefit function of sorters based on the coal gangue image information database to achieve optimal allocation of multiple tasks for multiple sorters, and to integrate reinforcement learning methods to achieve intelligent collaborative control of multiple sorters and optimal configuration of the number of sorters. This can effectively improve the sorting efficiency of the multiple-sorter system. Following the academic ideology of “Sorting gangue is sorting images”, our team independently developed a double robotic arm truss type coal gangue sorting robot experimental platform, which verified the correctness and feasibility of this academic ideology and successfully applied it in Yuhua Coal Mine of Tongchuan Mining Company. The academic ideology that “Sorting the gangue is sorting the images” has laid a theoretical foundation for solving the problems of intelligent, precise, and efficient coal gangue sorting.

LiDAR SLAM faces challenges in the narrow and confined unstructured environment of underground coal mines, where inaccurate point cloud pose estimation due to few or complex features can result in distortion or even map construction failure. To address the difficulties in LiDAR point cloud feature extraction and registration in this degraded environment, a two-stage method integrating FPFH and ICP algorithms is proposed. Initially, the method constructs kd-tree structures for the source and target point clouds, reduces point cloud numbers through statistical and voxel filtering, extracts point cloud surface normal, and computes fast point feature histogram descriptors for key points. Subsequently, a coarse registration is performed using the sampling consistency initial registration algorithm, followed by fine registration using the ICP algorithm to enhance point cloud registration accuracy and pose estimation precision. Furthermore, enhancements are made to the feature extraction and registration algorithm of the LIO-SAM, along with the optimization algorithm of the back-end loopback factor, to improve key local feature identification and registration capabilities. The addition of the Scan Context global descriptor loop factor enhances loop detection accuracy for consistent global mapping. Experimental testing on the M2DGR public dataset and SLAM experiments in simulated coal mine scenarios demonstrate the effectiveness of the improved algorithm in feature extraction and registration of the point clouds. Compared to the traditional LIO-SAM algorithm, the improved algorithm showcases higher accuracy in pose estimation and point cloud registration, with a 6.52% improvement in average relative position error and an 18.84% reduction in maximum absolute position error. The resulting maps exhibit no obvious distortion and mapping errors are within 1%, allowing for the construction of high-precision consistent global maps in unstructured and degraded environments.

The fusion positioning of high-precision fiber-optic inertial navigation and high-precision position sensor is an effective method to realize the accurate positioning of coal mine roadway roadheader. However, the high-precision fiber-optic inertial navigation has high cost and the error accumulates with time. How to achieve high-precision fiber-optic inertial navigation performance and eliminate cumulative error through low-cost and low-precision fiber-optic inertial navigation adaptive zero-speed correction is an urgent problem to be solved. Therefore, an adaptive zero-speed correction method for fiber-optic inertial navigation of coal mine roadheader based on zero-speed detection and extended Kalman filter is proposed.Aiming at the problem of inaccurate zero-speed detection of traditional threshold method for roadheader fiber-optic inertial navigation, a zero-speed detection method based on PCA−SCSO−SVM ( Principal Component Analysis PCA, Sand Cat Swarm Optimization SCSO, Support Vector Machine SVM ) is proposed. This method uses roadheader vibration signal for zero-speed detection. Firstly, the vibration signal is decomposed by VMD and the IMF component is selected according to the correlation coefficient. Secondly, the time-frequency domain features of IMF components are extracted, and the principal component analysis method is used to reduce the dimension to reduce the complexity of the diagnostic model and the difficulty of data analysis. Finally, the accuracy of zero-speed detection is improved by introducing the sandcat swarm optimization algorithm to optimize the kernel function and penalty parameters.Aiming at the problem of high cost and error accumulation with time of high-precision fiber-optic inertial navigation, an adaptive zero-speed correction method is proposed. According to the correction interval time determined by the zero-speed detection results of the roadheader and the motion characteristics of the roadheader, the speed error and angular velocity error of the extended Kalman filter at the zero-speed moment are used as observations to perform adaptive zero-speed correction. In order to verify the effectiveness of the proposed method, the experimental verification of zero-speed detection and zero-speed correction is carried out. In the zero-speed detection experiment, this method, SVM method, GA−SVM method and PSO−SVM method are compared. The experimental results show that the zero-speed detection accuracy of this method is the highest, reaching 96.5%.The zero-speed correction experimental results show that the zero-speed correction method proposed in this paper can effectively reduce the attitude error of the fiber-optic inertial navigation and improve the attitude detection accuracy of the roadheader. The shorter the correction interval, the more accurate the error estimation and the higher the corrected attitude accuracy. When the correction interval is 10 minutes, the fiber-optic inertial navigation of 0.1（°）/ h can reach the attitude detection accuracy of 0.057（°）/h, and the low-precision fiber-optic inertial navigation can reach the high-precision positioning target.

In the application of visual SLAM(Simultaneous Localization and Mapping) in coal mines, lighting changes and low-texture scenes seriously affect the extraction and matching of feature points, resulting in the failure of pose estimation and affecting the positioning accuracy. Therefore, a binocular vision localization algorithm SL-SLAM for underground mobile robots in coal mines based on the improved ORB (Oriented Fast and Rotated Brief)-SLAM3 algorithm is proposed. For the lighting change scenario, the original ORB feature point extraction algorithm is replaced by the SuperPoint feature point extraction network with lighting stability at the front end, and a feature point grid limitation method is proposed to effectively eliminate the invalid feature point area and increase the stability of pose estimation. For the low-texture scene, a stable LSD(Line Segment Detector) line feature extraction algorithm is introduced at the front end, and a point-line joint algorithm is proposed, which groups the line features according to the feature point grid, and matches the line features according to the matching results of the feature points, so as to reduce the matching complexity of the line features and save the pose estimation time. The reprojection error model of point features and line features is constructed, the angle constraints are added to the line feature residual model, the Jacobian matrix of the pose increment of point features and line features is derived, the unified cost function of the reprojection error of point features and line features is established, the local mapping thread uses the ORB-SLAM3 classic local optimization method to adjust the pose of points, line features and keyframes, and performs loop correction, subgraph fusion and global BA(Bundle Adjustment) in the back-end thread. The experimental results on the EuRoC dataset show that the APE(Absolute Pose Error) index of SL-SLAM is better than other comparison algorithms, and the trajectory prediction results closest to the true value are obtained, and the root mean square error is reduced by 17.3% compared with ORB-SLAM3. The experimental results of simulating the underground scene of coal mine show that SL-SLAM can adapt to the scene of light change and low texture, and can meet the positioning accuracy and stability of the mobile robot in the underground coal mine.

After coal mine disasters, the environment deteriorates and rescue tasks are complex and arduous, resulting in numerous threats to the rescue personnel. Robot participation in rescue can effectively improve rescue efficiency and safety, but existing remote sensing rescue robots suffer from problems, e.g., inability to communicate in real-time. Therefore, in response to the complex post-disaster downhole environment, a multi-robot autonomous exploration method based on the extended undirected graph is proposed, aiming at studying the collaborative search and rescue of a multi-robot autonomous exploration system to further improve rescue efficiency. Firstly, based on the characteristics of the downhole environment and the needs of post-disaster rescue, the architecture and algorithm flow of the multi-robot autonomous exploration system for coal mine rescue are constructed by combining local and global planning strategies to address the issues of computational efficiency and spatial exploration depth in autonomous exploration. Secondly, when conducting local rescue exploration, there is a special environment where narrow and open spaces coexist, making the multi-robot system difficult to quickly sample and generate local exploration paths. This situation can result in premature termination of rescue tasks. Therefore, the traveling salesman problem is integrated, where each robot performs viewpoint sampling and exploration gain calculation in the local space to construct a local map and determine the points to be visited. The A * algorithm is used to optimize the shortest exploration path in the local map. Moreover, when the local graph exploration gain is insufficient, global graph exploration is performed. Each robot shares the global maps and expands them incrementally to reduce the overall efficiency decline caused by multiple robots directly accessing the target point. The collaborative global map search algorithm solves the shortest path for each robot's global map exploration. Finally, three different models of robots are used for real experiments and compared with commonly used multi-robot autonomous exploration algorithms based on boundary point methods. The results show that the multi-robot autonomous exploration method proposed in this paper improves exploration completeness by 51% and saves exploration time by more than 58%. The proposed method can effectively achieve multi-robot autonomous exploration tasks in complex environments.

Under the background of intelligent mine construction, the application of intelligent equipment has increasingly become the main content of mine intelligent transformation. Intelligent robots in coal mine that are designed for inspecting and dangerous area surveying and doing other tasks depends on the construction of digital map of underground coal mine and the localizing of the robot itself. But most of the traditional localizing methods are inefficient or even ineffective in the underground. Simultaneous Localization and Mapping (SLAM) has become a better choice for underground intelligent robot localization methods. However, due to the high cost of lidar and the poor performance of camera in low illumination environment, it is necessary to design a SLAM localization method that takes into account both low cost and adaptability to low illumination environment. Therefore, a localization method of robot with underground dark light environment adaptability in coal mine is proposed. Firstly, the real images of the gallery of a coal mine in Fengxian County, Baoji City, Shaanxi Province and the dataset of the camera and IMU required for SLAM were collected. According to the images, non-matching dark light and normal light dataset was made, and 3560 images were obtained after data amplification. An EnlightenGAN image enhancement network combined with self-attention module is designed, which takes into account the dependence of different regions of the image without relying on the paired dataset. Based on the ORB–SLAM3 framework, the whole local image detection is introduced to screen the input image, and an improved IMU initialization strategy based on analytical solution is introduced to improve the initialization speed, and the improved image enhancement network is transplanted to enhance the low illumination and uneven illumination images. Experiments on the EuRoC dataset show that the image enhancement-based underground coal mine robot localization method can reduce E_RMS by 12.17% and E_SD by 14.35% in low-light environments. In two actual coal mine roadway scenarios, the low-light environment can be identified, and the increasing number of feature points are extracted and the drift phenomenon of positioning trajectory is reduced by the SLAM system and. Finally, the localizing effect of the system is improved in the dark area of the roadway.

Fully mechanized caving is the main mining method for extra-thick coal seams in my country. The control accuracy of the caving actuator depends largely on the data feedback of the tail beam posture. In order to improve the control accuracy of the caving actuator, a method for predicting the inclination of the support tail beam based on the long short-term memory neural network (LSTM) was proposed. The absolute coordinates of the support bottom plate, the inclination of the tail beam, the relative height of the tail beam, the frame shifting rate and the column pressure related to the tail beam caving action were used as the input variables of the RNN convolutional network and the LSTM neural network. The historical data of coal caving in a fully mechanized caving working face of a coal mine were used to train and verify the support tail beam posture prediction model, and the support tail beam posture prediction model was established. The tail beam inclination was predicted for 16 consecutive hours. The fitting degree of the predicted tail beam inclination curve and the actual tail beam inclination curve reached 98.7%. In the fully mechanized top coal caving face, 3～4 production shifts of hydraulic support tail beam inclination prediction tests were carried out. After comparing and analyzing the predicted tail beam inclination curve with the actual tail beam inclination curve, when the confidence interval was (0.98, 1.02), the prediction accuracy for 16 hours of continuous production was 98.40%. The LSTM-based support tail beam inclination posture prediction method solved the problem of tail beam inclination control in the electro-hydraulic control system's adaptive coal caving operation, laying the foundation for unmanned coal caving in the fully mechanized top coal caving face.

This paper presents the development of a megawatt-scale intelligent fracturing pump system for underground coal mines, designed to address the escalating demands for flow and pressure in large-scale regional fracturing applications, particularly in hard roof management and enhanced gas extraction permeability. The system integrates automatic control and variable frequency technology with the design of underground coal mine fracturing pumps, enabling dynamic collection of performance parameter data at various stages of hydraulic fracturing. It provides real-time analysis of the power matching for electrically driven fracturing pumps and achieves full automation of the fracturing process. Key technological challenges, such as the development of special materials for high-pressure, high-flow-rate fracturing pumps; the reliability of the transmission system and hydraulic ends for megawatt-level fracturing pumps; and intelligent control technology, have been successfully addressed. The research included: Development of high-strength, erosion-resistant martensitic precipitation hardening stainless steel suitable for extreme conditions with large flow, ultra-high pressure, and sand-mixed media, along with ultra-high pressure self-reinforcing treatment to enhance the fatigue life of the hydraulic ends of fracturing pumps; Investigation of critical reliability technologies, including high-strength welding for alloy steel, high-load-bearing, high-power-to-weight ratio gear transmission technology, and wear-resistant friction pairs of aluminum bronze alloy-cast iron, to ensure the reliability of the transmission system under high-power conditions; Creation of a high-durability metal plunger-combination seal fracturing fluid sealing pair, with the application of computer simulation technologies such as virtual prototyping, FEA, CFD, and hydraulic system simulation to optimize the structure, performance, and reliability of the fracturing pump's fluid end suction and discharge systems; Mastery of technologies such as low-frequency variable flow sealing for deep boreholes, automatic identification of coal and rock layer fracturing, and cyclic fracturing control, enabling intelligent control throughout the fracturing process. The industrial trial of this system has been successfully conducted at the Caojiatan coal mine 122110 extra-thick coal mine working face for weak zone management of hard roof strata and at the Dongli coal mine 1250 gas control lane for high-efficiency extraction of anti-reflection gas in coal seam areas. Field tests demonstrated that at the Caojiatan coal mine, pre-splitting treatment for hard roofs achieved stable fracture expansion pressure with a maximum of 32.4 MPa and an average flow rate of 100 m³/h. At the Dongli Coal Mine, the gas permeability enhancement test revealed that after 10 days of hydraulic fracturing, the average pure gas extraction volume increased to 1.596 m³/min, approximately 29 times that of the conventional drilling extraction process.

Expanding the speed range of high water-based piston pump is beneficial to the wide intelligent control of liquid supply system in fully mechanized coal mine. Aiming at the speed limitation of high water-based piston pumps caused by insufficient lubrication of key friction pairs under traditional low-pressure lubrication conditions, the group developed a high water-based piston pump structure and its synchronous lubrication method based on a stepped plunger to meet the high-speed operation. In order to deeply understand the system characteristics and synchronous lubrication effect of the friction pair, a system co-simulation model based on the software AMESim/Simulink is built to study the dynamic coupling characteristics between the lubrication and the working cavity, as well as the oil film characteristics of the slipper pair, which can reflect the lubrication effect of the friction pair under synchronous lubrication conditions. Further, the influence of pump speed on it is discussed. Finally, the flow characteristics of high water-based piston pump at high speed were tested by prototype test. The results show that the proposed synchronous lubrication method can produce lubricating oil in the lubrication chamber with increasing speed. With the increase of pump speed, the thickness of slipper auxiliary film decreases, and the extension of static pressure action time causes the increase of leakage. The oil film thickness of slipper boots ranges from 8.1 to 17.5 μm at the speed range of 500 to 2 000 r/min, meeting the needs of fluid lubrication. The experimental results show that when the speed is 500 r/min and 1 500 r/min, the deviation of the simulated and experimental average output flow of the piston pump is about 7.8% and 8.1%, respectively. The flow pulsation increases with the increase of the speed, and the measured flow pulsation is smaller than the numerical calculation result. The novel high water-based piston pump meets the demands of high-speed operation. Meanwhile, the establishment of pressure in the piston working chamber and lubrication chamber, as well as the opening and closing of the suction and discharge valves will have a hysteresis relative to the plunger movement, and become manifest with the increased pump speed, thus affecting the system volumetric efficiency and flow rate of pulsation. The research results verify the feasibility of the new synchronous lubrication high water-based piston pump applied to the high-speed conditions, and the causes of existing problems are analyzed, which will lay a theoretical and experimental basis for subsequent improvement and optimization.

The preparation and characterization of coal-based carbon materials is a hot topic. The structural characteristics of coal-based graphene not only determine the performance of the products, but also provide reference for process control during product preparation. Therefore, it is necessary to have a clear understanding of the microstructure and morphology of coal-based graphene. This study collected 8 different coal ranks coals in Ningxia, and obtained coal-based graphite through high-temperature graphitization treatment after chemical demineralization. Graphene was prepared using an improved Hummer’s redox method, and the microstructure and structural parameters of coal-based graphene were comprehensively characterized by X-ray diffraction, Raman spectroscopy, atomic force microscopy, and high-resolution transmission electron microscopy. The results indicate that the coal–based graphene products prepared from coal samples of different coal ranks in Ningxia are all few-layer graphene. Graphitization can significantly improve the order and graphitization degree of coal structure, which is beneficial for the preparation of coal-based graphene, especially for low coals. In this study, the graphitization degree of coal-based graphites are more than 0.70, and the highest is 0.99. The higher the coal rank, the larger the diameter and stacking height of the graphene prepared, the fewer defects, and the better the peeling effect. and better the stripping effect of graphene prepared. The diameter of the coal-based graphene prepared in this study ranges from 3.78 to 13.04 nm, with interlayer spacing exceeding 0.355 0 nm and a maximum of 0.366 1 nm. As the coal rank increases, the diameter of coal-based graphene increases more significantly than the stacking height, the morphology of coal-based graphene tends to develop in a flattened form in general. The caking of coals can cause an increase in the surface roughness of graphene, but it does not affect the development of the crystal form and order of graphene microcrystalline domains. As the coal rank increases, the graphene structure undergoes an evolution path of vortex type -Y-type - semi-plane type -plane type. The plane type graphene structure is the final form of coal structure evolution during the preparation of coal-based graphene, characterized by obvious peeling, curling, and stretching, which also provides the possibility for the formation of step like of graphene structures.

To investigate the desulfurization effect and mechanism of fly ash as a desulfurized, single factor experiment and response surface method were used to analyze the parameters of ash slurry solid-liquid mass ratio , SO₂ volume fraction, and gas flow rate, and the reaction mechanism of fly ash wet desulfurization was investigated by X-ray fluorescence spectroscopy and scanning electron microscopy. The results show that within a certain range, increasing the solid-liquid ratio can increase the penetration time, the total adsorption capacity of SO₂, and the adsorption capacity of SO₂ per unit of fly ash slurry by increasing the pH. However, when the solid-liquid ratio exceeds 1∶1, the penetration time and adsorption capacity of SO₂ will decrease. With the increase in SO₂ volume fraction, the adsorption capacity of fly ash slurry to SO₂ first increased and then decreased. When the SO₂ volume fraction was 750×10⁻⁶, the maximum adsorption capacity was 64.35 mg, High SO₂ volume fraction can significantly reduce the penetration time. With the increase of the gas flow rate, the penetration time, the total adsorption amount of SO₂, and the adsorption amount of SO₂ per unit mass of fly ash slurry decreased. The results of response surface analysis are shown by denoting the three factors of solid-liquid ratio, SO₂ concentration and gas flow rate as A, B and C, respectively, and their interaction terms as AB, AC and BC, respectively: the primary and secondary order of the factors affecting the penetration time were C, B, A, BC, AC, and AB. The primary and secondary order of the factors affecting the total adsorption capacity of SO₂ is B, A, C, AB, BC and AC. The primary and secondary order of factors affecting the adsorption of SO₂ per unit mass of fly ash slurry is A, B, C, AB, AC and BC. The optimum desulphurisation process conditions are as follows: the solid/liquid ratio is 0.87, the SO₂ volume fraction is 472×10⁻⁶ and the gas flow rate is 1 500 mL/min. XRF and SEM analysis of fly ash before and after desulfurization show that the sulfur oxide content of fly ash after desulfurization increases significantly, and the original surface plate Ca(OH)₂ forms CaSO₄·2H₂O and CaO·Al₂O₃·3CaSO₄·32H₂O in bulk and rod. The possible desulfurization mechanism is that fly ash dissolves a large amount of Ca²⁺ and OH⁻ in water. SiO₂ and Al₂O₃ react with Ca(OH)₂ under strong alkaline conditions to form cementitious materials such as hydrated calcium silicate and hydrated calcium aluminate (C-S-H and C-A-H gels).When the sulfur-containing flue gas passes into the fly ash slurry, SO₂ is transferred to the liquid phase, forming H⁺, ${\mathrm{HSO}}_3^-$ and ${\mathrm{SO}}_3^{2-}$, and H⁺ reacts with the fly ash slurry, leaching Ca²⁺, Fe³⁺ and other elements contained in the fly ash and catalyzing oxidation of SO₂ gas dissolved in the fly ash slurry to produce H₂SO₄ and CaSO₄. After combining with water molecules in the slurry, CaSO₄ was precipitated in the form of CaSO₄·2H₂O.

Coal, as an abundant carbon resource, holds a crucial position in the energy industry. However, conventional coal utilization methods are linked to environmental pollution and inefficiency. At the molecular level, coal and its derivatives display properties of polycyclic aromatic hydrocarbons, which could be transformed into valuable low-dimensional carbon materials (0D, 1D, and 2D) through physical and chemical processes. This transformation plays a significant role in achieving cleaner coal utilization and promoting the development of renewable energy sources. The article first elaborates on the current situation of energy and environment in China, as well as the urgent need for clean utilization of coal resources in the “dual carbon” era. The structural features, preparation methodologies, modification strategies, and new energy applications of coal-derived low-dimensional carbon materials, including carbon quantum dots, carbon nanotubes, graphene, carbon nanosheets, and etc. are explored. Additionally, it elucidates the benefits and limitations, innovation, as well as suitable conditions of various preparation techniques of low-dimensional carbon materials. Furthermore, it investigates the diverse applications of these materials in the components of organic solar cells (OSC), highlighting their application superiority and mechanisms as electrodes, additives, and charge transport layers in detail. The pivotal role of coal-based low-dimensional carbon materials in enhancing the power conversion efficiency of OSC, reducing manufacturing costs, improve stability, and supporting sustainable development is analyzed. The application advantages and performance indicators of low-dimensional carbon materials compared to other low-dimensional materials in OSC are discussed. Finally, the problems and key research directions in the preparation and new energy application research of coal-based low-dimensional carbon materials were analyzed and summarized. With the continuous development of high-value carbon material preparation technology, the reduction of production costs, and large-scale production, these materials are anticipated to have even broader application prospects in the realm of renewable energy.

In order to reveal the structural characteristics of microbial communities in acid mine drainage(AMD) under different redox environmental conditions and the interaction between microbial communities and groundwater environment, Select the typical abandoned coal mining area in the Shandi River basin of the Niangziguan Spring area, and collect AMD samples from different redox environments for Hydrochemical characteristics and isotope analysis and the high-throughput sequencing of the V4, V5 region of microbial 16S rRNA. The analysis of Hydrochemical characteristics and isotope indicates that the hydrochemical characteristics of AMD are mainly influenced by pH and surrounding rock. The analysis of hydrochemical isotopes shows that the hydrochemical characteristics of AMD are mainly influenced by pH and rock water interaction. Under aerobic and anaerobic environments, the pH of AMD is negatively correlated with Eh; The pH of AMD in anaerobic and hypoxic environments is negatively correlated with Fe³⁺and Eh.The Ca and Mg ions in AMD mainly come from the dissolution of calcite and dolomite in the surrounding rock and coal. The K and Na ions in AMD mainly come from the dissolution of silicate rocks such as feldspar under acidic conditions; The main source of ${\mathrm{SO}}_4^{2-}$ and Fe content in AMD is the oxidation of pyrite in coal-bearing strata; The sulfur isotope evolution of AMD is characterized by lower δ³⁴S–${\mathrm{SO}}_4^{2-}$ values in aerobic environments and higher δ³⁴S–${\mathrm{SO}}_4^{2-}$ values in anaerobic and hypoxic environments. The analysis of the microbial diversity impact of AMD in the bottom of the mountain basin shows that the lower the pH, the lower the bacterial abundance and diversity of the sample; The higher the temperature, the higher the bacterial abundance and diversity of the sample. Through the correlation analysis between the microbial community and environmental factors of AMD, it was found that the dominant bacterial genera in the Liugou are Ferrovum, Gallionella and Ferritrophicum. The dominant genera of AMD in the Miaogou open-pit mine are Ferrovum, Gallionella and Acidithiobacillus. The dominant genera of AMD in Yulinnao drilling are Ferrovum and Gallionella. The microbial community of AMD in the Shandi River Basin is greatly affected by pH, temperature and precipitation.