Numerical simulations have often been used in close-distance coal seam studies. However, numerical simulations can contain certain subjective and objective limitations, such as high randomness and excessively simplified models. In this study, close-distance coal seams were mechanically modeled based on the half-plane theory. An analytical solution of the floor stress distribution was derived and visualized using Mathematica software. The principal stress difference was regarded as a stability criterion for the rock surrounding the roadway. Then, the evolution laws of the floor principal stress difference under different factors that influence stability were further examined. Finally, stability control measures for the rock surrounding the roadway in the lower coal seam were proposed. The results indicated the following: (1) The principal stress difference of the floor considers the centerline of the upper coal pillar as a symmetry axis and transmits radially downward. The principal stress difference in the rock surrounding the roadway gradually decreases as the distance from the upper coal pillar increases and can be ranked in the following order: left rib > roof > right rib. (2) The minimum principal stress difference zones are located at the center of the left and right “spirals,” which are obliquely below the edge of the upper coal pillar. This is an ideal position for the lower coal seam roadway. (3) The shallowness of the roadway, a small stress concentration coefficient, high level of coal cohesion, large coal internal friction angle, and appropriate lengthening of the working face of the upper coal seam are conducive to the stability of the lower coal seam roadway. (4) Through bolt (cable) support, borehole pressure relief, and pregrouting measures, the roof-to-floor and rib-to-rib convergence of the 13313 return airway is significantly reduced, and the stability of the rock surrounding the roadway is substantially improved. This research provides a theoretical basis and field experience for stabilizing the lower coal seam roadways in close-distance coal seams.

Because both faults and cleats exist in coal, sharp stress drops occur during loading when coal is deformed. These drops occur during the pre-peak stage and are accompanied by sudden energy releases. After a stress drop, the stress climbs slowly following a zigzag path and the energy accumulated during the pre-peak stage is unstable. A stress–strain curve is the basic tool used to evaluate the bursting liability of coal. Based on energy accumulation in an unsteady state, the pre-peak stress–strain curve is divided into three stages: pre-extreme, stress drop, and re-rising stage. The energy evolution of the specimen during each stage is analyzed. In this paper, an index called the effective elastic strain energy release rate (EESERR) index is proposed and used to evaluate the coal’s bursting liability. The paper shows that the propagation and coalescence of cracks is accompanied by energy release. The stress climb following a zigzag path prolongs the plastic deformation stage. This causes a significant difference between the work done by a hydraulic press during a laboratory uniaxial compression experiment and the elastic strain energy stored in the specimen during the experiment, so the evaluation result of the burst energy index would be too high. The determination of bursting liability is a comprehensive evaluation of the elastic strain energy accumulated in coal that is released when the specimen is damaged. The index proposed in this paper fully integrates the energy evolution of coal samples being damaged by loading, the amount of elastic strain energy released during the sample failure divided by the failure time is the energy release rate. The calculation method is simplified so that the uniaxial compressive strength and elastic modulus are included which makes the new index more universal and comprehensive. Theoretical analysis and physical compression experiments validate the reliability of the evaluation.

Mining above confined aquifer has become an important task for water inrush prevention in China. To study the failure characteristics of stope floor along the strike, a mechanical model under combined action of mining and confined aquifer was constructed, and the distribution of vertical stress, horizontal stress, and shear stress was obtained. Based on the Mohr–Coulomb criterion, the failure range of the floor is determined and verified by the in situ test. The results indicate the following. (1) Both vertical stress and horizontal stress in the stope floor take the junction of stress increasing area and stress decreasing area as the dividing line, forming two groups of “convex arches” at the solid coal side and the goaf side, respectively. (2) The vertical stress gradient in the solid coal side is significantly higher than that in the goaf side, while the horizontal stress gradient in the solid coal side is similar to that in the goaf side. The shear stress distribution is divided into three regions by the boundary between positive and negative shear stress, which makes the stope floor in this area to show compression shear or tension shear failure. (3) According to the in situ test, the maximum floor failure depth of 41503 working face is 11.38 m, which is quite close to the theoretical calculation result of 9.68 m. (4) Applying the mechanical model to five other coal mines with different mining conditions and stress states, the maximum absolute error between the measured and theoretical values of floor failure depth is 1.1 m, the average absolute error is 0.8 m, the maximum relative error is 8.2%, and the average relative error is 6.5%. The study provides a certain mechanical basis and reference for the floor failure mechanism induced by mining and confined aquifer.

Rock burst hazards are a common occurrence during deep coal mining, resulting in the collapse of roadway wall rock and the failure of support structures. In response, in situ testing was conducted in underground mine areas affected by rock burst disasters to systematically assess the impact of rock burst on wall rock of roadways and the performance of support systems. The findings of the study suggest that rock burst events lead to a substantial decrease in the strength of the coal mass of the roadway roof and ribs, as well as an increase in fractured zones. Additionally, there is a reduction in wave velocity, attenuation of anchoring and pretension forces of rock bolts, and varying degrees of damage to anchoring interfaces between rock bolts and their anchoring agents and between anchoring agents and wall rock. The parameters defining the mechanical and impact resistance properties of rock bolts and anchor cables, including extension rate and tensile strength, exhibit a significant decrease. In light of the aforementioned research results, it is recommended that the existing support system for roadways vulnerable to rock burst be enhanced and reconfigured in three key areas: improving support design, choosing materials with superior impact resistance capabilities, and bolstering monitoring and warning mechanisms.