To address the problem of pillar instability in residual mining areas under repeated mining disturbances, the prediction of pillar instability time and mining technology is studied by using numerical simulations, theoretical examinations, and practical experimentation techniques utilizing the remaining facade of the Shenghua Coal Industry 3103. On the basis of the findings, employing a fully mechanized mining technique exerts greater pressure on the coal pillar, increasing the likelihood of plastic failure. However, adopting the segmental filling mining approach can effectively disperse this pressure, consequently mitigating potential damage to the coal pillar. Initially, the coal pillar undergoes a progressive yielding or collapsing process at its periphery, leading to instability once the area of the projectile core of the coal pillar decreases beyond a specific threshold. The dimensions of both the affected roadway and the coal pillar play a pivotal role in determining the coal pillar's load‐bearing capacity. A prediction model for pillar instability is established. If the area of the projectile core is less than 34% during remining, the mining method should be changed or pillar reinforcement should be implemented. The accurate prediction and adaptation of the model's mining approach ensure the secure extraction of coal pillars within the remaining mining region, with successful validation of the model's performance.

This study examines the failure laws of soft and hard interlayer overlying rocks and the development height of the water‐conducting fracture zone in a gently inclined super‐thick coal seam at the Qifeng coal mine, Shanxi. The overlying rocks above the working face were classified into four typical structural types: soft‐hard‐soft‐hard, hard‐soft‐hard‐soft, soft‐soft‐hard‐hard, and hard‐hard‐soft‐soft. First, the failure laws of these four types of overlying rock structures were studied theoretically. Theoretical analysis revealed that different overlying rock structures form different composite rock beams that move synchronously after coal excavation. Numerical simulation using FLAC^3D showed that the overburden failure does not change linearly but forms different combined rock beams according to the different overburden structures, resulting in a step change. The soft–soft–hard–hard overburden structure exhibits the largest damage extent, while the soft–hard–soft–hard overburden structure results in the highest damage height. When the hard rock above the coal seam is closer to the stope, the influence of the mining stress is greater, leading to stress concentration at the junction of soft and hard rocks. Three methods—theoretical analysis, numerical simulation, and field measurement—were used to study the overlying rock failure law and the development height of the water‐conducting fracture zone of the 51408 working face of the Qifeng Coal industry. The results show that the guiding height of the 51408 working face of the Qifeng Coal industry is 120.12, 103.6, 113.2, and 116.1 m for the four overlying rock structures. The maximum measured height of 116.1 m was considered the development height of the water‐conducting fracture zone. The field‐measured results are consistent with those obtained by theoretical calculation and numerical simulation, confirming the accuracy of the theoretical analysis and numerical simulation and providing a basis for water prevention and control of the roof of gently tilting super‐thick coal seams.

With the increasing depth of engineering in rock masses, the issue of significant rheological deformation becomes notably prominent. To accurately characterize the deformation and failure behavior of the surrounding rock near the goaf in deep small coal pillars, we have developed a viscoelastic–plastic constitutive model that takes into account the initial stress associated with creep. This model builds upon an existing viscoelastic–plastic constitutive framework and is implemented numerically using C++ in the FLAC^3D. The Oedometer test method is used to check the creep components and the calculated parameter (η/G) ensure convergence of numerical operations. We conducted both pre‐improvement and post‐improvement tests of the constitutive model in a simple numerical analysis of roadways. The enhanced model demonstrates improved accuracy in deviational stress calculations and proves suitable for modeling deep, soft roadways. Furthermore, we applied the improved model to simulate actual working conditions at Zhaozhuang Mine, where mining operations near the working face induce a fracture zone of approximately 3 m in the small coal pillar located at the goaf's edge. The new model closely aligns with the observed results, further validating its suitability for tunneling along the goaf. Tunneling and mining activities result in significant deformation at various locations, including the shoulder socket of the working face, the top angle of the coal pillar, and the bottom angle of the floor. The crushing zone at the working face extends up to 4 m. Rheological effects transform the triangular elastic core, from its original high‐pressure elastic state to a fracture and plastic zone in the deep coal. Stress in deep coal is transmitted along the triangular elastic core. Numerical simulation results closely match the excavation profile of the tunneling, providing valuable insights into the impact of mining activities. This study can provide reference for deformation of similar geological conditions and engineering situation.

Under the influence of engineering disturbances, the loading rate of surrounding rock is in a state of continuous adjustment. This study conducts experimental investigations on the mechanical response characteristics under different strain rates (10⁻⁵ s⁻¹, 10⁻⁴ s⁻¹, and 10⁻³ s⁻¹). During the uniaxial loading process of coal–rock composite specimens, multi-parameter monitoring was implemented, and a systematic study was carried out on the ring-down count induced by microcracks, the energy values of acoustic emission (AE) events, the stage-dependent strain characteristics on the specimen surface, and the surface temperature variation characteristics. Additionally, the stress–strain curve characteristics under different strain rates were comparatively analyzed in stages. The loading process of the coal–rock composite specimens was reproduced using the Particle Flow Code (PFC3D 6.0) simulation software. The simulation results indicate that the stress–strain results obtained from the simulation are in good agreement with the laboratory test results; based on these simulation results, the energy accumulation and dissipation characteristics of the coal–rock composite specimens under the influence of strain rate were revealed. Furthermore, a microscopic damage model considering strain rate was constructed based on the Weibull probability statistics theory. The results show that strain rate has a significant impact on the strength, elastic modulus, and failure mode of the coal–rock composite specimens. At low strain rates, the specimens exhibit obvious progressive failure characteristics and strain localization phenomena, while at higher strain rates, they show brittle sudden failure characteristics. Meanwhile, the thermal imaging results reveal that at high strain rates, the overall temperature rise in the composite specimens is rapid, whereas at low strain rates, the overall temperature rise is slow—but the temperature rise in the coal portion is faster than that in the rock portion. The peak temperature at high strain rates is approximately 2 °C higher than that at low strain rates. The PFC simulation results demonstrate that the larger the strain rate, the faster the growth rate of plastic energy in the post-peak stage and the faster the release rate of elastic energy.

This study elaborates on the concepts of total elastic energy, elastic volumetric energy, elastic deviatoric energy, total plastic energy, plastic volumetric energy, and plastic deviatoric energy involved in the deformation and failure of rock masses. A plastic energy increment formulation is further derived. Based on in‐situ measurements of the maximum, intermediate, and minimum principal stresses in the Xiaojiawa Coal Mine, representative stress values are substituted into the theoretical model, enabling an accurate characterization of the energy evolution of the roadway surrounding rock under the influence of mining‐induced approaching excavation. Numerical simulations are subsequently conducted to analyze the energy evolution and deterioration behavior of the surrounding rock during the approaching process, and the results agree well with the theoretical predictions. During the advance of the 120803 longwall panel toward the 221303 belt roadway, the surrounding rock at 2.5 m into the solid‐coal rib remains in a persistent plastic state, indicating pronounced degradation. When the distance between the 221303 material roadway and the 120803 working face reduces to approximately 100 m, the plastic zone initiates and extends from the solid‐coal side at a depth of ~2.5 m toward greater depths, accompanied by a progressive deepening of the surrounding‐rock deterioration zone. Concurrently, the elastic strain energy around the roadway gradually decreases, whereas the plastic dissipation energy continues to increase. When the belt roadway is 100 m ahead of the longwall face, the surrounding rock at 6.5 m into the solid‐coal rib begins to undergo a transition from elastic energy accumulation to plastic energy dissipation, signifying severe and irreversible deterioration. The plastic zone has extended approximately 6.5 m into the solid coal side. In contrast, the rock mass at 10.5 m remains elastic. When the roadway and longwall face become aligned, elastic energy begins to accumulate at 10.5 m, and the deterioration zone expands correspondingly.