The drilling pressure relief technology is an effective way to reduce the accumulation of elastic energy in the tunnel envelope, which can reduce the risk of regional ground pressure occurrence. However, there is a lack of theoretical guidance on which drilling parameter has the greatest degree of influence on the effectiveness of pressure relief. The uniaxial compression tests were conducted to study the relationships between drilling parameters (the diameter, depth, and spacing) and the mechanical properties and deformation modulus of specimens. The results show that: (1) The drilling diameter (DDR) and drilling depth (DDH) of single-hole specimens negatively correlate with the peak-failure strength and deformation modulus, while the drilling spacing (DS) of double-hole specimens positively correlates with the peak-failure strength and deformation modulus. It shows that the borehole diameter has a more significant effect on the decompression effect. (2) With the help of the Grey Relational Analysis, the factors affecting the peak-failure strength and deformation modulus of the drilled specimens were ranked in significance. From the largest to the smallest, they are DDR, followed by DDH and DS. (3) The role of the pressure relief mechanism is to transfer the high stress in the shallow part of the roadway to the deep part, reduce the peak strength of destruction and deformation modulus of the peripheral rock in the drilled section, so that the characteristics of the mechanical behavior of the rock are significantly weakened, and the range of the area of the drilled hole decompression is enlarged. During the loading of the borehole, the borehole stress field dominates in the early stage, and cracking starts near the borehole along the direction perpendicular to the direction of maximum principal stress (horizontal direction). In the later stage, the maximum principal stress field dominates and vertical cracks with large widths appear. During crack expansion, the plastic energy dissipation effect is enhanced and the deep impact conduction path is weakened, thus protecting the roadway. This study determined the significance of the pressure relief effect of different drilling parameters, which can guide reasonable modifications of drilling parameters in the field.

One of the main challenges in excavating roadways is implementing temporary supports that are powered by hydraulics and have high strength. The current temporary support system lacks active support and often causes separation between the top plate and the layer below. It is crucial to control the initial separation of the roadway roof for the stability of the surrounding rock, especially on roadways with loose and soft rock. This research focuses on the A4027 return airway in Sail Six Mine. The issues with the temporary support system in this airway have been identified. The concept and principle of using hydraulically driven, high-strength temporary support technology are proposed. A mechanical analysis model is created to study the stacked roof in the temporary support region, and the critical conditions for delamination of the top plate are determined. The relationship between the delamination difficulty parameter Q, the distance between temporary supports L, and the strength of the temporary supports q is quantified. Numerical simulation using Flac3d is used to model the relationship between the strength of the temporary supports and the deformation and stress of the rock on the roof. The overall strength of the temporary supports for the A4027 return airway is determined to be 10 kN/m², with a distance of 2 m between the temporary supports. Hydraulically driven, high-strength temporary support devices are developed and tested for their strength. Field trials are conducted as well. The results show that the initial separation of the top plate is improved and that the support effect in the temporary support region is significant. The maximum separation of the top plate during excavation is only 34 mm, and the sinking of the top plate does not exceed 68 mm. This effectively limits the deformation of the surrounding rocks in the very soft coal seam, providing valuable insights for other roadways with similar conditions.

To study the effect of mining dynamic response characteristics on the footwall working face of the normal fault under the influence of the gob area, theoretical research, indoor experiment, and numerical simulation are adopted to analyze the stress manifestation characteristics, overburden movement, and energy evolution characteristics during the process of mining. The results show that: (1) In the process of mining toward the fault, the working face shows the change characteristics of “stable-activation mutation-final stability”. At 20 m from the fault, the arch structure of the working face was damaged, fissures appeared near the high fault fracture zone, and the displacement of the overburden rock increased significantly; (2) the maximum value was reached at 4–8 m from the coal wall, and the superposition of tectonic stress and mining stress led to the concentration of the stress and energy accumulating on the top plate near the fault, and the data close to the gob area were even larger; (3) If the plastic damage zone of the high-level rock layer on the hanging wall and footwall of the fault appears to have a wide range of penetration, and the area formed between the shear displacement curve of the fault plane and the X-axis appears to have a significant enhancement, it is considered that the fault has been activated; (4) The size of the coal pillar of the fault is determined to be 40 m, and combined with the pressure unloading technique of the variable-diameter drilling hole, the validation is carried out through the micro-vibration monitoring, and the results of which can be used as a reference for the safety of the working face under similar conditions.

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.