Rockburst of deep roadway was induced by the superposition of mine earthquake disturbance and high static stress exceeding the limit strength of coal‐rock mass. To study the roadway impact instability characteristics caused by mine earthquake disturbance and to propose an optimized support scheme, the discrete element model of the roadway structure was established based on the 1305 working face of the Zhaolou Coal Mine. The influence of mine earthquake amplitude and hypocenter location on the roadway was analyzed. The mesocrack evolution characteristics of the roadway were simulated and reproduced. Characteristics of stress field, crack field, displacement field, and energy field of the disturbed roadway with different support schemes were studied. The results showed that the greater the amplitude of the mine earthquake was, the severer the roadway impact failure was. The upper and left hypocenters had a significant influence on the roadway. The superposition of the high static stress and the dynamic stress due to the far‐field mine earthquake resulted in the impact instability of coal‐rock mass around the roadway, causing severe roof subsidence as well as rib and bottom heave. The evolution of tensile cracks caused the severe impact failure of roadway from a mesoscopic perspective. Using the flexible support to reinforce the roadway retarded the stress decline in roof and rib, improved the self‐stability, reduced the number of near‐field cracks, and decreased the displacement. Meanwhile, it allowed the roof and rib deformation, which was conducive to releasing elastic energy in surrounding rocks and reducing mine earthquake energy. The cracks and deformation in the floor were controlled by using the floor bolt. The optimal support scheme for a roadway to resist mine earthquake disturbance was proposed: “bolt‐cable‐mesh‐steel strip‐π‐beam + floor bolt.” The research results have a specific guiding significance for the support of the coal mine roadway.

In order to explore the mechanism of floor dynamic rupture, the current study adopts a thin plate model to further investigate the condition of floor failure. One of the possible explanations could be floor buckling due to high horizontal stress and dynamic disturbance ultimately leading to rapid and massive release of elastic energy thus inducing dynamic rupture. Seismic computed tomography and 3D location were employed to explore the evolution characteristics of floor stress distribution and positions of mine tremors. In the regions of floor dynamic rupture, higher P‐wave velocity was recorded prior to the dynamic rupture. On the contrary, relatively lower reading was observed after the dynamic rupture thus depicting a high stress concentration condition. Meanwhile, evolution of mine tremors revealed the accumulation and subsequent release of energy during the dynamic rupture process. It was further revealed that dynamic rupture was induced due to the superposition of static and dynamic stresses: (i) the high static stress concentration due to frontal and lateral abutment stress from coal pillar and (ii) dynamic stress from the fracture and caving of coal pillar, hard roof, and key stratum. In the later part of this study, the floor dynamic rupture occurrence process would be reproduced through numerical simulations within a 0.6 sec time frame. The above‐mentioned findings would be used to propose a feasible mechanism for prewarning and prevention of floor dynamic rupture using seismic computed tomography and mine tremors 3D location.

In order to explore the mechanism of rockburst in coal seam with rock parting, a combination of on-site and numerical experiment is used to study the failure and instability process, crack propagation mechanism, and influencing factors. The following four points were addressed: (1) the instability is a process that roadway in coal seam with rock parting go through from stable locking in the initial stress unloading stage to slipping unlocking, and then to spatter ejection in slipping dynamic load disturbance stage. (2) The fracture development caused by unloading excavation of coal seam with rock parting will change from shear crack to tensile crack. In this process, coal-rock contact surface slip and coal-rock fracture are coupled with each other. (3) The greater the mining depth is, the greater the lateral pressure coefficient is, and the higher the rockburst risk is. On the contrary, the lower the risk of rockburst. (4) When choosing the support form of roadway in coal seam with rock parting, the two supporting forms of bolting (cable) and supplementary masonry support should be preferred. The results enrich the theory of the dynamics of surrounding rock fracture in coal mine, further clarify the potential dangers to mining-area roadways and working faces, and provide technical information to ensure the safe and efficient mining of bifurcated coal seam.

Exploring the mechanical behavior of deep rocks, determining the development laws of microcracks, and enriching the evaluation indicators of rock damage are prerequisites for ensuring the safety and stability of underground engineering. In this research, by carrying out triaxial experiments and based on mechanical analysis, acoustic emission (AE) damage characteristics study and Three-Dimensional Particle Flow Code numerical simulation, it is determined that confining pressure can limit internal crack propagation, delay the formation of macroscopic failure surface, reduce damage accumulation rate, and strengthen rock mechanical properties. The development of microcracks inside rocks presents a process of “local aggregation of microcracks to form damage area, continuous development of microcracks to expand damage area, penetration of damage zone to form fracture zone, penetration of fracture zone to form macroscopic failure surface”, which is the fundamental cause of rock damage, and the fracture zone formed by the accumulation of AE high-energy events is the main controlling factor for the formation of rock macroscopic failure surfaces. As quantitative evaluation indicators of rock damage status, AE S value and AE Q_t value are introduced, which can correspond to various stages of rock crack development. The research results can provide some benefits for the safety and stability of deep underground engineering.

An underground coal gasification (UCG) process is strongly exothermic, which will cause thermal damage on rock cap. We proposed a new thermal damage numerical model based on a two dimension particle flow code (PFC2D) to analyze the inception and extension of cracks on pre-cracked red sandstone, which were thermally treated at a temperature of 25~1000 °C. The results indicated that: (1) a thermal damage value D_T obtained by extracting the thermal crack area of scanning electron microscope (SEM), which can be used as an indicator of the degree of thermal damage of the sandstone; (2) a thermal damage numerical model established by replacing the flat-joint model with the smooth-joint model based on the thermal damage value D_T, this approach can properly simulate the mechanical behavior and failure patterns of sandstone; (3) the critical temperature for strength reduction was 750 °C. The peak strength increased as pre-treatment temperature increased from 25 to 750 °C and then decreased. The elastic modulus E₁ decreased with the increasing thermal treatment temperature; (4) micro-scale cracks initiate from the tip of the prefabricated fissure, and expand along the direction of prefabricated fissure, finally developing into macroscopic fracture. This approach has the potential to enhance the predictive capability of modeling and presents a reliable model to simulate the mechanical behavior of thermally damaged sandstones, thereby offering a sound scientific basis for the utilization of space after UCG.