Coal mining has gradually entered the deep mining era, and large-height mining is an important way to mine thick coal seams in the deep. The high coal wall will inevitably make the distribution of the overburden structure in the coal mining face more complicated, and the large buried depth will also cause more intense mine pressure. The study of the distribution and evolution of the overburden structure and stress in the mining site can provide theoretical guidance for safe mining. In this work, a physical similarity modeling test was carried out based on the physical–mechanical parameters of overburden rock and similarity theory, taking the mining of a deep, large-height working face in Pingdingshan Coal Mine as an example. The results show that the deformation and breakage of overburden rock in deep, large-height workings occurring during mining is persistent and not only in a short period of time. The breakage form of overburden can be categorized into two types based on the deformation characteristics: (I) non-separation-induced type, and (II) separation-induced type. Among these, the breakage induced by separation can be divided into two categories: (i) dominated by self-weight stress, and (ii) affected by shear cracks. It also summarizes the form of the overburden structure and the structural morphology of the stope. The overburden structure shows a “combined cantilever beam structure-articulated rock-slab structure-non-articulated rock-slab structure”. Among these, the periodic breakage of the upper cantilever beam evolved articulated and non-articulated rock-slab structure in the lower part, which weakened the supporting effect of the lower gangue and further aggravated the breakage of the upper overburden rock. The shape of the main structure of the stope mainly depends on the fracture line from the advancing coal wall to the upper overburden: from a rectangular shape without collapse to a trapezoidal shape at the initial stage of collapse, to a trapezoidal shape with multiple steps after the main roof collapse.

Due to the influence of the ground stress, mining disturbance, and other factors, the roadway surrounding rock in deep underground engineering such as mines, tunnels, and underground caverns is prone to looseness and deformation with the excavation of roadways. In such engineering, the bolt support is frequently employed to stabilize the surrounding rock. In this work, a part of the anchor and the surrounding rock were taken as a simplified model of the anchorage rock mass, and the laboratory compression test was performed on the similitude model. Then, the FLAC3D software was used to simulate varying numbers of bolts and different lateral pressure conditions, and the peak stress, the maximum principal stress field, and the anchor stress field distribution of the anchorage rock mass were obtained. The influence of bolt pretightening force and row spacing on the stability of surrounding rock was discussed using the combined arch theory. The results show that increasing the number of bolts and lateral pressure in the anchorage rock mass can significantly improve the stress value and distribution range of the maximum principal stress field and the anchorage stress field. The fluctuation of the anchorage stress field at different anchorage distances can be lessened by increasing the number of bolts (bolt density). When the lateral pressure exceeds 3 MPa, the anchorage mechanical characteristics of the anchorage rock mass tend to remain stable. The coverage of the effective anchorage stress field and the thickness of the surrounding rock anchorage composite arch can be increased by increasing the bolt pretightening force and decreasing row spacing, consequently improving the anchorage mechanical characteristics of the anchorage rock mass. The research results can be used as a theoretical reference for choosing appropriate bolt support parameters for the roadway surrounding rock.

Anchored surrounding rock is prone to large nonlinear deformation and instability failure under dynamic disturbances. The fissures and defects within the surrounding rock make the rock mass’s bearing characteristics and deformation instability behavior increasingly complex. To investigate the effect of the fissure angle on the dynamic mechanical response of the anchored body, a dynamic loading model of the anchored, fissured surrounding rock unit body was established based on the finite difference–discrete element coupling method. The main conclusions are as follows: Compared to the indoor test results, this numerical model can accurately simulate the dynamic response characteristics of the unit body. As the fissure angle increased, the dynamic strength, failure strain, and dynamic elastic modulus of the specimen generally decreased and then increased, with a critical angle at approximately 45°. Compared to 0°, when the fissure angle was 45°, the dynamic strength, failure strain, and dynamic elastic modulus decreased by 17.08%, 15.48%, and 9.11%, respectively. Additionally, the evolution process of cracks and fragments shows that the larger the fissure angle, the more likely cracks are to develop along the initial fissure direction, which then triggers the formation of tensile cracks in other regions. Increasing the fissure angle causes the specimen to rupture earlier, making the main rupture plane more directional.