Large deformation control of deep soft rock roadways has been a major problem in mining activities worldwide. This paper considers the supporting problem related to large deformation of a deep soft rock roadway in Chao’hua coal mine. The discrete element simulation method (UDEC software) is adopted to simulate a tailgate of panel 31041 in Chao’hua coal mine. The failure patterns of unsupported and primary supported roadway are simulated, and these reveal the characteristics of deformation, stress and crack propagation. The excavation of roadway leads to high deviator stress, which exceeds the peak strength of shallow surrounding rock and causes it to enter the post-failure stage. Tensile failures then initiate and develop around the roadway, which causes the fragmentation, dilation and separation of shallow surrounding rock. The compressive capacity of the primary support system is low, which results in serious contraction in the full section of the roadway. An improved control scheme is put forward for the support of a tailgate. The underground test results confirm that the improved support system effectively controlled large deformation of the surrounding rocks, which can provide references for support in the design of roadways excavated in deep soft stratum.

In order to reduce large deformation failure occurrences in non-pillar longwall mining entries due to roof weighting behaviors, a case study in Halagou coal mine was conducted on optimization and control techniques for entry stability in non-pillar longwall mining. The Universal Discrete Element Code (UDEC) modeling was adopted to study entry stability in non-pillar mining, and the characteristics of deformation and stress and crack propagation were revealed. The large deformation transmission between the entry-immediate roof and the gob-immediate roof could be eliminated by optimizing the entry roof structure through a directional roof-cutting method. The localized tensile stresses generated in the entry-surrounding rock caused the generation of coalescent macroscopic fractures, which resulted in the instability of the entry. The tensile stress state could be inhibited by an active flexible support system through enhancing the confining pressure on the surrounding rock. Serious rotation subsidence occurs in the entry roof due to periodic weighting of the main roof, which could be greatly reduced by a passive rigid support pattern. The numerical and field test results both showed that the roof weighting pressure was offloaded by the technique and that the deformation of the entry surrounding the rock in non-pillar mining was quite small. Thus, the technique can effectively ensure the stability of the gob-side entry, which can provide references for entry stability control in non-pillar longwall mining.

Uniaxial compression experiments were carried out on a combination of 12 different lithology and different coal-rock height ratios to study the effects of lithology and rock-coal height ratio on the mechanical properties and impact effects of the composite. The mechanical constitutive models of coal, rock-coal-rock combination were used to analyze the instability process of the combined body and the evolution of physical parameters such as displacement, acceleration, stiffness, energy of various parts during the loading process. The criterion for the stability and instability of the combined body with the stiffness of post-peak of coal and the stiffness of rock unloading as the basic parameter was proposed. The results show that with the decrease of rock-coal height ratio, the failure types of the combined body are “clastic” complete failure, “Y-type” semi-complete failure, “local” incomplete failure (“clastic”, “Y-type” and “local” refer to the failure form of the specimens, complete failure, semi-complete failure, and incomplete failure refer to the failure degree of the specimens), and the compressive strength and elastic modulus of the combined body are reduced, and the impact energy index of the combined body is increased; the harder the rock components in the combined body are, the greater the compressive strength, elastic modulus and impact energy index of the combined body are; the failure state of the combined body is determined by the post-peak stiffness λ of the coal, and the contrast relationship between the stiffness and the rock stiffnesses k₁, k₂ of the roof and floor. After the peak is unloaded to the point S where the change rate of the stiffness of the coal curve meets ${\lambda }^{\prime }\ge k_1{k}_2\ddot{m}/\left[\left(k_1+k_2\right){\dot{u}}^2\right]$, the combined body changes from stable failure to unstable failure. The release of the roof and floor during the failure of the combined body is proposed. The ratio α of the energy released by the roof and floor to the coal consumption energy is used as an evaluation index of outburst proneness under coal and rock combination, and the criterion is proposed.

In order to solve the problems of the high cost and time consumption of half coal rock entry driving, low coal recovery rate, and stress concentration on filling support body of retained entry along gob, the innovative 110 mining method based on pressure relief by roof cutting was adopted in 6302 thin coal seam working face of Baoshan Coal Mine. First the technical principle and key technology of this mining method was presented. Then, through theoretical analysis and calculation, engineering experience, and field test, the key parameters such as the length of constant resistance anchor cable, the cutting angle and height of presplitting blasting, the charge structure, and the blocking‐gangue support structure were determined and conducted in the retained entry. The broken expanded coefficient varying law of caved gangue with time and space was obtained, which revealed roof movement characteristic. The displacement monitoring curve of the roof and floor indicated that the maximum subsidence of the roof was about 150 mm and the maximum amount of floor heaving was 100 mm, which were quite small. The field monitoring data indicated that the entry retaining effect is good, which indicated that the innovative 110 mining method can be an effective way for reducing the high cost and time consumption of half coal rock entry driving, enhancing the coal recovery rate and preventing the dynamic mine pressure disasters.

After mining, upward fractures develop from bottom to top, and they are key channel for groundwater and gas flow and also sand burst; therefore, it is important to study the development mechanism and criterion of upward fractures. Combining with physical simulation and theoretical analysis, the development law of mining‐induced upward fractures was revealed, and the mechanism and criterion of fracture propagation was studied. The results show that the development of mining‐induced upward fractures can be divided into the following three stages: with the face advances from the open cut, the vertical fractures are generated at the mining boundary, and the roof is flexural but not caved; it is “Fractures generate stage.” After roof caves, the upward fractures extend and develop upward along the caving angle with continue advances; it is “Fractures develop and extend stage.” When the face reaches critical mining, the development height of the upward fractures is at its maximum and is basically invariable; it is “Fractures stabilization stage.” Mining‐induced fractures can be analyzed as mixed‐type fractures in fracture mechanics; when the combined stress exceeds the critical strength of stratum, it will extend downward until penetrate through the stratum, and the upward fracture develops and extends upwards. The parameter controlling the fracture of the rock stratum is the maximum hoop tensile stress σ_(θ)max, at the fracture end, the theoretical model of fracture propagation was established, and the criterion of upward fractures was proposed.