The application of roof-cutting and pressure-relief gob-side entry retention plays a critical role in controlling the stability of the surrounding rock at the entry, easing continuity tension and improving resource recovery ratio. The excavation of the 360,803 airway in Xinji No. 1 Mine is affected by intense mining of the 360,805 working face. Hence, to address the stability problem of surrounding rock in the 360,803 airway, rock mass blast weakening theory was used in this study to analyze the blasting stress of columnar charged rock mass and obtain the radiuses of crushed, fractured, and vibration zones under uncoupled charging conditions. The reasonable array pitch, length, and dip angle of boreholes were determined according to the pressure-relief range of the blasting fracture. The migration laws of roof strata were explored based on a mechanical model of overlying roof strata structure on the working face. Subsequently, the horizon, breaking span, and caving sequence of hard roof strata were obtained to determine the roof-cutting height of this entry. On the basis of the theory of key stratum, the number of sequences at the roof caving limit stratum and hanging roof length in the goaf were calculated, the analytical solution to critical coal pillar width was acquired, the evaluation indexes for the stability of entry-protecting coal pillars were determined, and the engineering requirements for the 25 m entry-protecting coal pillars in the 360,803 airway were met. Moreover, various indexes such as roof separation fracture, displacement of surrounding rock, and loose circle of surrounding rock in the gob-side entry were analyzed. The stability and cementation status of surrounding rock in the 360,803 airway were evaluated, and tunneling safety was ensured.

Rock excavation has experienced complex stress paths. The development of the original crack under the path of principal stress magnitude and principal stress direction is a key scientific problem that needs to be solved in rock underground engineering. The principal stress magnitude dominates the initiation and propagation of the crack and increases rock damage under the action of principal stress rotation. In this study, the theoretical calculation and numerical analysis method have been combined with the crack propagation conditions to study the stress‐driven mechanism of brittle rock crack propagation under principal stress rotation. The results show that the “relative initial angle” of crack angle is being updated in time during the principal stress rotation process; once the stress is rotated, it will become the next initial crack angle; the crack propagation direction is deviated under the applied shear load, and it is always in the direction of minimum shear load, leading to a certain degree of inhibition of crack propagation depth in the initial direction. According to the results of numerical simulation, the effect of principal stress rotation caused by mining excavation is obvious and has a certain range of influence depth, the stress of surrounding rock of roadway is the highest within the depth range of 1∼2 m, and the maximum principal stress is as high as 26.89 MPa. The rotation of principal stress direction on the roadway surrounding rock surface is the strongest, which makes the surrounding rock more fragmented, and the middle principal stress and the maximum principal stress rotate about 90° counterclockwise along the Ox axis. Studying the action mechanism of principal stress rotation on fractured rock masses can provide scientific basis for geotechnical engineering design and rock mass surrounding support.

An overlying rock structure plays a key role in controlling the roof deformation of nonpillar gob‐side entry retaining by roof cutting. On the bases of the actual geological conditions of II 632 Haulage Roadway at the Hengyuan coal mine, a similar three‐dimensional simulation experiment of roof precutting is conducted. Thereafter, the caving characteristics and migration law of the roof strata in the strike and dip directions are obtained. Moreover, the roof of the retained roadway and key strata of the goaf can form a hinge structure of the key blocks. By monitoring the deformation of the surrounding rock and stress distribution of the roof, the skew deformation characteristics of roadway roof are obtained. By observing the borehole peeping technology, the roof subsidence near the goaf is determined to be greater than that of the solid coal side, and the roof subsidence of the gob‐side entry retained by roof cutting is greater than that of the floor heave and two sides approaching. Results of the three‐dimensional similar simulation experiment indicate that the mechanical structure model of the key block of the retained roadway roof is constructed, and the mechanical analytical solution of the required support resistance of the retained roadway roof is obtained. This study proposes the constant resistance and large deformation anchor cable reinforcement support method to control the roof deformation of the retaining roadway. Through engineering application, the maximum value of the roof and floor movement of the retained roadway is stable at approximately 650 mm. The retained roadway can meet the demand of the next mining face.

The 360,606 working face of Xinji 1^# Mine was taken as an engineering case to address the significant safety hazards associated with coal mining under hard roof. A key technical scheme for pre-splitting blasting roof-cutting pressure relief was developed by theoretical analysis, numerical simulation, and field experiments. The results are as follows. (1) Fracture propagation was dominated by synergistic effects of explosion gases and stress waves. Explosive energy was concentrated at crack tips after exploration. When accumulated energy exceeded the rock’s deformation capacity, cracks propagated further. (2) Explosive energy distribution was optimal between the crushing zone and fracture zone at a decoupling coefficient of 1.5, which avoided energy waste from over-fragmentation. A 5-m spacing yielded a peak effective stress of 98.6 MPa at the midpoint between boreholes. Integrated stress cloud analysis indicated that a spacing of 5 m was optimal. (3) Post-implementation displacements were measured as follows: roof (238 mm), coal pillar sidewall (124 mm), mining sidewall (100 mm), and floor (48 mm). The successful retention of the 306,606 working face gateway validated the efficacy of this technology in roof-cutting and roadway protection. These findings provide technical guidance for coal extraction under analogous engineering.

After coal extraction in underground mining, progressive fracturing occurs in both roof and floor strata, resulting in stress redistribution within these rock layers. Considering the fractures in the strata at roof and floor, an elliptical stress arch was introduced. This concept analyzed how the shape and structure of this arch evolved when mining progressed, as well as the evolution law of the strata at roof and floor fractured. Research results are as follows. (1) The elliptical stress arch bore most of the load on the surrounding rocks of the mining site, and the front and rear arch springing moved forward with the mining face. The surrounding rocks of the mining site redistributed stress, which formed tensile stress in the areas near the arch top and base. The stratum underwent active fracture and was compressed in its movement direction. Compressive stress was formed to clamp unbroken strata in the area far away from the arch top and base. Strata were passively broken along the arch trajectory due to their self-weight load, the compressive load of active-fracture strata, and the clamping effect of the compressive stress around the stress arch. (2) Unbroken hard strata inside the elliptical stress arch formed a cantilever beam. A hinged rock beam was formed after hard strata broke. Soft strata acted on lower strata in the form of loads. The support structures at the working face and the coal wall carried the weight of strata, which was transmitted through cantilever beams or hinged rock beams at roof and floor. The passive fracture of the cantilever beam and the instability of the hinged rock beam directly affected the stability of the working face support and coal wall. (3) The analysis focused on how the lengths of the long and short axes and the coordinates of the center of the elliptical stress arch changed as the working face advanced. The theory of composite beams was used to determine strata, rock loads, fracture distances, and fracture sequences of hard strata at the roof and floor of the elliptical stress arch. The stability at the roof and floor of the hinged rock beam was obtained using the criteria for compression deformation instability and sliding instability of the hinged rock beam. The research results can provide support for calculating the resistance of the working face support, the load borne by the coal wall, and the width of the plastic zone of the coal wall. Besides, they offer a theoretical basis for analyzing the stress of the working face support and the stability of the coal wall in the mining area, as well as production practice.

The stress state of primary rocks is redistributed during coal mining, which leads to deformation and failure of the floor strata. Coal seam floor experiences repeated additional stress from the mining of two overlying seams in close-distance mining. An assessment of floor stability is required in the presence of a confined aquifer within the floor strata. The failure characteristics of floor rock strata are crucial for analyzing water inrush from floor. This mechanical response of floor deformation and failure becomes particularly pronounced when the roof is treated using the full caving method. Floor strata were analyzed for the 220,105 working face at Xinji 2^# Coal Mine, with an elastic half‑plane theory established. The stress distribution characteristics of the floor were investigated using software MATLAB. The width of the plastic zone and the maximum failure depth of floor were determined based on the slip line field theory of floor rock mass. A reference was provided for water damage control during repeated mining of the 1^# coal seam. Continuous field observation of floor failure was performed through drilling combined with the network parallel electrical method. Apparent resistivity at the depth of 19.6 m below floor was approximately twice the background resistivity of the corresponding stratum. The apparent resistivity of rock strata below 19.6 m remained constant, indicating a protective zone. These results confirm the reliability of the theoretical analysis and provide a technical foundation for safe mining in the working face.