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In view of the problems associated with the poor stability of coal walls, coal slide and leakage of top-coal at the tunnel excavation working face under a soft and extra-thick coal roof, the surrounding rock at the tunnel excavation working face must be strengthened. The theoretical analysis of rock pressure, numerical simulation and other methods were comprehensively used to study the coal-wall-slicing mechanism. Given the characteristics of a soft and extra-thick coal roof, the combined supporting technology of “coal wall water injection + metal roof frame” is proposed. The findings show that in the process of roadway excavation, the coal–rock junctions of the wall and the middle part of the roof are weak areas that are prone to spalling and therefore need to be strengthened. Laboratory tests determined the moisture content of the coal body during tunneling to provide data for the parameter design of coal wall water injection. Safe and efficient excavation of the roadway was ensured by injecting water into the coal wall in combination with a metal roof protection skeleton. The application of this technology not only effectively prevents rib spalling but improves control of the deformation of the surrounding rock. During 40 days of field observation, the maximum deformation of the roof was 24.8 mm, and the distance between the two roadway walls was 21.5 mm. The deformation of the roadway was controlled within a safety zone. The application of this technology reduced the repair rate of the roadway and improved the efficiency of the roadway excavation. It brought significant economic benefits and provides an important reference for similar mines.

To study the mechanism of rockburst and its spatio-temporal evolution criterion, a rockburst simulation experiment was performed on granite specimens, each with a prefabricated circular hole, under different lateral loads. Using micro camera, acoustic emission (AE) system, and infrared thermal imager, the AE characteristics and thermal radiation temperature migration were studied during the rockburst process. Then, the failure mode and damage evolution of the surrounding rock were analyzed. The results demonstrate that increasing the lateral load can first increase and then reduce the bearing capacity of the hole. In this experiment, the hole failure process could be divided into four periods: quiet, particle ejection, stability failure and collapse. Correspondingly, the AE signals evolved from a calm stage, to have intermittent appearance; then, they were continuous with a sudden increase, and finally increased dramatically. The failure of the surrounding rock was mainly tensile failure, while shear failure tended to first increase and then decrease. Meanwhile, damage to the hole increased gradually during the particle ejection period, whereas damage to the rockburst mainly occurred in the stability failure period. The thermal radiation temperature migration exhibited warming in shallow parts, inward expansion, cooling in the shallow parts with free surface heating, inward expansion, a sudden rise in temperature of the rockburst pits, and finally specimen failure. The initial reinforcement support should fully contribute to surface support. Furthermore, an appropriate tensile capacity and good energy absorption capacity should be established in support systems for high-stress roadways.

During the gob-side entry driving of the Jurassic coal seam in Western China, ejection disasters occur under the influence of roof drainage, which rarely appear in the eastern mining area. To address this problem, a method combining theoretical analysis, numerical simulation, and field monitoring was used to study the disaster induced by head-on ejection during speedy driving under the influence of roof drainage in the context of gob-side entry driving of the 2202 auxiliary haulage roadway in a mine. A calculation model for the critical energy conditions for disasters induced by head-on ejection was established. The relationships between the driving velocity and the dynamic and static loads on the driving face and the disasters induced by ejection were clarified under the influence of roof drainage. The results indicate that the energy threshold for ejection-induced disaster is 12.23 kJ, and the elastic energy of the driving face induced only by static load fails to reach the energy threshold. When the driving velocity exceeds 5 m/d, microseismic activity in front of the driving face increases in a stepwise manner, and the influence of dynamic load intensifies. The superposition of accumulated elastic energy induced by static load and the energy transmitted by microearthquake to the driving face exceeds the energy threshold, resulting in the risk of ejection-induced disaster. Based on this, measures such as microseismic monitoring in front of the driving face, advanced pulverized coal monitoring, driving speed adjustment, and advanced pressure relief of large-diameter boreholes are adopted to ensure safe driving.

Anthracite in a specific area of Shanxi Province is the subject of this essay’s research. In the creep studies, different porosity intervals and pore water pressures were employed to evaluate the mechanical properties of creep under various test paths. The conventional Burges model was coupled in series with the nonlinear viscous elements and plastic elements. The key parameters in the equation are fitted, and a creep model is created to describe the nonlinear viscosity-elastic-plastic characteristics of coal under the influence of pore water pressure with varying porosities. The creep tests used varied porosity intervals, pore water pressures, and test paths to study the mechanical properties of creep. The conventional Burges model was coupled in series with the nonlinear viscous element and plastic element. To represent the nonlinear viscosity-elastic-plastic properties of coal under the effect of pore water pressure with variable porosities, the main parameters in the equation are fitted, and a creep model is developed. The results show that the porosity and strength of the coal sample are negatively correlated. In comparison to coal samples with a porosity of 5–10%, the uniaxial compressive strength of coal samples with a porosity of 10–15% and 15–20% reduced by 9.6% and 22.3%. Throughout the creep process, instantaneous strain rises with porosity, and changes in pore water pressure and porosity have an effect on instantaneous creep under low-stress loading, resulting in different creep curve starting strain values. The duration from stress loading to the accelerated creep stage in the failure stage and the time from the deceleration creep stage to the accelerated creep stage are both gradually shortened with an increase in porosity and pore water pressure. For regression analysis and parameter identification, a creep constitutive model was developed to describe the creep characteristics of coal samples with varying porosity under varying pore water pressure. The creep parameters of the new constitutive model were obtained, and they could very well reflect the creep characteristics of specimens with varying porosity intervals under the influence of pore water pressure.

The problem needs to be solved about stability control of small coal pillars along goaf in the coal mining process. In this paper, the effects of water reducer, accelerator and expansion agent on the fluidity, setting time, expansion ratio and stone body strength of grouting materials were analyzed through orthogonal experiments, and the optimal ratio of grouting materials was obtained, and it is applied to the engineering site to obtain the reinforcement effect of the coal pillar grouting material on the coal pillar. The results of the study show that: When the dosage of the accelerating agent is 4%, the amount of water-reducer is 0.3%, and the amount of the expansive agent is 6%, the comprehensive performance of the grouting material is the best. After using the new coal pillar grouting material to strengthen the coal pillars, 30 days of monitoring were conducted to determine the deformation of the surrounding rock of the roadway. The shallow separation layer of the surrounding rock of the roadway was about 39.6~52.5 mm, and the accumulated separation layer of the deep rock was 28.5~29.5 mm, which were kept within the safe control range. The coal pillar grouting material can well fill the surrounding rock fissures and coupling and cement the broken rock mass, improve the bearing capacity of the narrow coal pillar, and enhance the overall stability of the surrounding rock of the roadway. The research results have important reference significance for similar coal pillar grouting.

The protective layer mining method of the traditional deep coal seam in has been confronted with great challenges, and it is difficult for coal and gas to be extracted together. Taking the occurrence conditions of III1 mining area of Luling Coal Mine located at Huaibei, China, as engineering background, the influence law of the lithology on stress environment in front of the stope was analyzed by theoretical analysis and numerical simulations. The mining-induced mechanical effect of coal-rock mass was studied under different protective layer mining modes. The results showed that the peak value of the advanced abutment pressure decreased with the decrease of lithologic strength under the same mining conditions. For simulated geological conditions, the stress concentration coefficient of soft rock and coal seam protective layer mining modes was 1.9 and 1.7, respectively. Under the mining stress path of different protective layers, the ratio of axial stress increase and confining pressure unloading in secondary unloading phase were 2 : 1 and 1.5 : 1, respectively. The axial stress-strain curves of different protective layer mining modes had similar trends, and they had a volume expansion at the end of unloading (failure stage). In addition, it revealed the pressure-relief antireflection mechanism of the protective layer mining. Under the same confining pressure condition, the peak stress and peak strain increased with the increase of loading and unloading velocity ratio. The reduced value of the confining pressure increased, while the volume expansion decreased at failure. The results were applied to III1 mining area in soft rock protective layer mining, which created the mining way of traditional coal seam protective layer. Furthermore, the gas control technology of soft rock protective layer working face was put forward for deep coal seam with low permeability and high gas, enriching the pressure-relief mining theory of protective layer.

In order to study the creep behavior of deep soft rock, gritstone was chosen as the research subject, and a rock triaxial rheometer (Rock 600-50) and acoustic emission (AE) system (SH-II) were used to carry out the grade unloading confining pressure creep test under a high-stress level. The test results showed that the lateral creep behavior of the gritstone was more prominent than the axial creep under the initial high confining pressure. Under the same confining pressure, the creep strain rate (the direction the same as strain) of the gritstone decreases with the increase in axial pressure. As shown by the AE count, AE signals were generated throughout the entire test process, indicating that the creep was a “microdynamic” process. The creep behavior was characterized by a significant confining pressure effect. As the confining pressure was decreased, the degree of creep increases significantly. During the test, the AE energy increased on the whole but decreases during the creep phase. During the entire test process, the overall energy in the constant deviatoric stress grade unloading of the confining pressure was 45% higher than that in the constant axial pressure grade unloading. The degree of failure of the rock was different in these two unloading creep tests, and the constant axial pressure grade unloading of the confining pressure entails greater damage than the constant deviatoric stress grade unloading of the confining pressure. The main reason was that the former had a lower confining pressure level and longer creep process than the latter, and the sample was mainly characterized by creep damage and large cumulative damage, while the latter features mainly unloading damage. Through the inversion of the Burgers constitutive model and nonlinear damage constitutive model for the creep test curve, the nonlinear constitutive equation can better fit the accelerated creep stage, which suggested that this model can describe the accelerated creep characteristics of the high-stress soft rock.

To deeply analyze the failure evolution of surrounding rock during excavation-induced unloading of the high-stress rock mass, a multistage failure model was established based on revealed failure patterns. The critical conditions for wing cracks were determined. The slab crack buckling analysis was carried out. The true-triaxial rockburst testing system was used for the miniature model test to study the fracturing evolution of surrounding rocks during excavation-induced unloading of the high-stress rock mass. The research results indicated that harder rock samples had higher compressive strength. Moreover, the smaller peak strains implied more obvious yield/plastic stages of harder rock samples with high confining pressures and softer rock samples with low confining pressures. V-shaped grooves appeared at the beginning of the surrounding rock’s failure while spalling and splitting occurred as the stress increased. Finally, the entire sample’s overall splitting failure was observed, and the borehole bottom bulged upward. The harder rock masses had fewer fractures and higher degrees of failure. There were obvious V-shaped grooves on both sides of the marble cave wall. The tensile failure occurred near the opening surface and shear failure at a far distance. The sandstone's overall failure was related to tensile cracking, and splitting failure occurred far away from the opening surface, which was similar to the in situ failure of surrounding rocks during excavation-induced unloading of the high-stress rock mass. The results obtained are instrumental in the construction safety control and prevention of underground engineering disasters.