Evaluations of the risk of fault-induced water inrush hazard is an important issue for mining engineering applications. According to the characteristics of the seam floor during mining advancing, a mechanical model of fault activation is built to obtain the equations of normal stress and shear stress on the surface of fault, as well as the mechanics criterion of fault activation. Furthermore, using FLAC^3D numerical software, the stress variation on the surface of fault under two different mining advancing directions are numerically simulated, and the distribution characteristics of the plastic failure zone of the roof and floor near the fault are obtained. The results show that: (1) When mining advances from the hanging wall, the normal stress increases more greatly than that from the foot wall, the shear stress distribution changes drastically with a large peak, and it is more likely to cause fault activation. (2) When mining advances from the hanging wall and approaches the fault, the normal stress and shear stress within the fault first increases, and then decreases suddenly. When mining advances from the foot wall, the normal stress and shear stress increases constantly, and the fault zone stays in the compaction state where the hanging wall and foot wall are squeezed together, which is unfavorable for water inrush hazard. (3) When mining advances from the hanging wall, the deep-seated fault under the floor is damaged first, and the plastic failure zone of the floor increases obviously. When mining advances from the foot wall, the shallow fault under the floor is damaged first, and the plastic failure zone of roof increases obviously. (4) For a water-conducting fault, the waterproof coal pillar size of the mining advancing from the hanging wall should be larger than that from the foot wall. (5) The in-situ monitoring results are in agreement with the simulation results, which proves the effectiveness of the simulation.

The damage and degradation of coal-measure sandstone in cold-region open-pit mines due to freeze–thaw effects has become one of the significant factors inducing instability in the rock mass of open-pit mine slopes. This study conducts experiments on the physical and mechanical properties of saturated coal-measure sandstone under varying freeze–thaw cycle counts and freezing temperatures, revealing the intrinsic mechanisms of damage and degradation in saturated coal-measure sandstone due to freeze–thaw effects. The experimental results indicate that, with an increase in the number of freeze–thaw cycles and a decrease in the freezing temperature, the elastic modulus and peak compressive strength of the specimens exhibit an exponential decrease. In contrast, the peak strain shows an exponential increase. However, compared to the freezing temperature, the increase in the freeze–thaw cycle frequency leads to a more significant change in the mechanical parameters of the specimens, indicating that the frequency of freeze–thaw cycles has a more pronounced effect on the deformation resistance of saturated coal-measure sandstone than the freezing temperature. The failure mode of coal-measure sandstone specimens under uniaxial compressive loading primarily exhibits shear failure; however, as the number of freeze–thaw cycles increases and the freezing temperature decreases, the specimens begin to exhibit tensile failure modes, which gradually develop into a combined tensile and shear failure mode. Based on the experimental data, two sets of surface equations were fitted to characterize the relationship between the mechanical properties (peak compressive strength, elastic modulus) of the specimens and the experimental parameters (number of freeze–thaw cycles, freezing temperature). The research findings can provide references and insights for engineering disasters caused by the degradation of coal-bearing sandstone in cold-region open-pit mines.

How to deal with hard rock cheaply and safely is a pressing issue in today’s coal mining. Weakening fractures of hard rock have always been a significant concern in China’s coal mine engineering. In this study, mechanical derivation, laboratory experiments, and numerical simulation research methodologies are used to evaluate the fracturing process of the static crushing agent (SCA). From a mechanical standpoint, the mechanism of fracturing hard rock by a crushing agent is investigated. It is assumed that single-hole fracturing is separated into three stages: the microfracture stage, the fissure development stage, and the breaking stage. The swelling and fracturing properties of SCA were quantitatively analyzed. It was found that the swelling pressure of SCA increased with the increase in pore diameter, and the range of the swelling pressure was 43.5 MPa to 75.1 MPa. SCA exhibited a delayed fracture initiation, but the rate of breakage was relatively high. The cracking effect of a single-hole specimen under no peripheral pressure was simulated using PFC2D, and the results were consistent with experimental observations. The internal dynamic effect, crack extension, distribution characteristics, and the development law of double-hole expansion pressure were analyzed for double-hole specimens with different hole diameters, hole spacings, and circumferential pressures. It was observed that the cracking effect was positively correlated with the pore diameter, while the pore spacing and surrounding pressure were negatively correlated. The size of the expansion pressure was negatively correlated with the pore diameter, while the pore spacing and surrounding pressure were positively correlated.

Insufficient stability of the top plate at the corner of an easily combustible coal seam comprehensive mining face may lead to a natural fire within the goaf. While corner sealing is crucial for minimizing air leakage, current sealing methods struggle to effectively prevent such leakage. Additionally, the distribution characteristics of the oxidation zone in the goaf after sealing are unclear, making it difficult to control the extent of the oxidation zone. To address these issues, a new type of inorganic paste filling material was developed, taking into account the conditions of the Cuncaota II Mine. Various corner-filling schemes were developed, and numerical simulations were used to study the effects of different corner-filling strategies and varying filling interval distances on the width of the oxidation zone in the goaf. Based on these findings, a working face corner-filling technology was proposed and applied to the 22,122 working face. The research results indicate that the mountain sand-based paste filling material, using mountain sand as the filling aggregate and cement and fly ash as the binding materials, not only meets the pumping requirements but also exhibits excellent self-supporting characteristics, thereby addressing the corner filling needs of the working face. The width variation in the oxidation zone in the goaf is influenced by the position and interval distance of the corner filling, showing a pattern of initially decreasing and then increasing with the rise in the filling interval distance, reaching a minimum at a filling interval of 50 m. Field observation data demonstrate that, following the application of the aforementioned filling technology, the width of the oxidation zone in the goaf of the 22,122 working face is reduced by 37.5%, and air leakage decreases by 66.7% compared to the unfilled condition. This technology effectively narrows the range of the oxidation zone in the goaf, ensuring the safety of working face production.

In coal mine roadway support, the arrangement of roadways in areas with high-stress concentrations is a challenging issue. It is crucial to solve this problem in order to ensure the safety and productivity of the mine. In this study, we selected a typical roadway in Zhong Xing Coal Mine, Fen Xi, Shanxi Province, as our research focus. We analyzed the deformation and failure characteristics of the surrounding rock under field mining pressure during roadway excavation through theoretical analysis, numerical simulation, and on-site measurements. Using the Flac3D numerical simulation technique, we studied the stress distribution characteristics of the roadway in the stress concentration area. To address the significant deformation of the roadway, we proposed a high-pretensioned top beam high-strength dual-anchor truss support system (HSCTS) and carefully studied its control mechanism and structural characteristics. After simulating and comparing its support effectiveness with the original support scheme and conducting on-site industrial tests, we determined the HSCTS support system as the alternative support system for roadway 1409. The test results showed that the convergence of the rib plates and the roof was no more than 79 mm and 191 mm, respectively. The significant influence of the roadway on the control of the surrounding rock also indicated effective control of the surrounding rock’s deformation along the 1409 goaf. The support control technology system can provide specific reference values for roadways under similar geological production conditions.

The stability criteria for surrounding rock in thick coal seam with multi-layer gangue roof are crucial theoretical foundations for assessing the risk of roof collapse in mine roadways. This study focuses on the deformation characteristics of surrounding rock in the thick coal seam with multi-layer gangue roof at the Tashan Mine, establishing a solid-supported beam mechanical model for the coal-rock layer combination. The mechanical criteria for shear failure of roof anchor cables and roof delamination are derived. Additionally, this paper employs 3DEC numerical simulations to investigate the distribution patterns of the plastic zone, stress field, and displacement field under varying lateral pressure coefficients.The research results show that the initial position of the plastic zone development of the roadway roof in the thick coal seam with gangue roof is the gangue layer of the roof, and further extension causes the failure of the supporting components until the instability of the roadway. When the lateral pressure coefficient λ > 1.2, delamination occurs in the shallow gangue layer of the roadway roof, resulting in a loss of load-bearing capacity. As the depth of the roof increases, the influence of the deep gangue layer on roof delamination gradually diminishes. Based on the engineering background of 5110 roadway in Tashan Mine, the control technology of ' roof bolting and grouting + grading support + anchor cable truss ( roof and sidewall ) ' is put forward. After the optimization scheme is adopted, the maximum deformation of roof is 178 mm, which is 61% lower than that of the original support, and the maximum separation amount is 92 mm. The bonding of the slurry with the roof strata forms a composite structure, enhancing the integrity of the surrounding rock and yielding favorable results in engineering applications.