After mining the upper-goaf side, large coal pillars and part of hard roof exposed above the pillars remain. The hard roof can significantly deform the roadway by transferring high stress through coal pillars to the roadway. This paper reports the use of hydraulic fracturing technology to cut the hard roof on both sides (i.e. the broken roof slides to the goaf) to relieve the pressure. The position of the roof fracture is the key to controlling the pressure relief. The bearing characteristics of the large coal pillars and hard roof are analyzed to establish a mechanical model of the broken-roof sliding instability after directional fracturing and determine the width of the coal pillars that can collapse under maximum overburden load on coal pillars as a reasonable hydraulic fracturing position. The results show that the distance from the mine gateway to the fracture location increases with increasing hard-roof length, coal pillar depth, coal pillar thickness (mining height), and goaf width. In addition, the distance to the mine gateway decreases with increasing coal strength, support of the coal pillar by the anchor rod, cohesive force, and internal friction angle of the coal–rock interface. Engineering tests were applied in coal roadway 5107 of coal seam 5# of the Baidong Coal Mine of the Datong Coal Mine Group. Given the site conditions, a reasonable fracturing length of 8.8 m was obtained. Next, directional hydraulic fracturing was implemented. The comparison of the roof deformation before and after fracturing suggests that this method reduces the local stress concentration in coal pillars, which allows the surrounding rock deformation in roadway 5107 to be controlled.

In order to study the development height of the water-conducting fracture zone in a fully mechanized top coal caving face. The E2311 working face of Gaohe Coal Mine was chosen as the research object, and the combined-strata structure and the rock layer synergistic movement mechanism were determined by combining engineering geological investigation and theoretical analysis. The height of the water-conducting fracture zone at the working face was calculated based on the combined-strata structure, and then the theoretical results were verified by numerical simulation and field measurement. The results show that after the coal seam is extracted from the working face, the movement of the overlying rock layers is in the form of bending and sinking movement of the rock layer group as a unit. Each rock layer group is controlled by a supportive lower layer with greater thickness and strength, driving the upper layers of weaker rock layers to synchronize and coordinate the movement; the sinking curvature is the same, after the lowermost support layer is bent and broken, its overlying weaker rock layers will move and break at the same time. The height of the water-conducting fracture zone of the working face were obtained by theoretical calculation, numerical simulation, and field measurements, which are 83.82 m, 84.3 m, and 86.6 m, respectively. The results are nearly consistent, thus the prediction of the height of the water-conducting fracture zone under the combined-strata structure is more accurate.

The bedding plane effect will occur when hydraulic fractures propagate to the bedding plane in sedimentary strata, resulting in the “≠,” “工,” or “/” shaped fracture morphology. Based on previous physical experiments results, this article analyzed the mecroscopic propagation mechanism of tensile failure and the mechanical conditions for main hydraulic fracture and the bedding plane fracture propagating, proposing the criteria for hydraulic fracture to penetrate through the bedding plane. A fully three-dimensional model of hydraulic fracture morphology in horizontal borehole hydraulic fracturing is established with the vertical water flow, water leak-off, and bedding plane effect taken into consideration. Basic equations of continuity, pressure decline, hydraulic fracture morphology, and others are solved. After that, true triaxial hydraulic fracturing experiments with samples containing bedding planes are conducted to verify the aperture, length, width, and height of hydraulic fractures in this model. The model is proved to be accurate and reliable.

In this paper, the gas–water two-phase seepage process under a real mechanical environment is restored by a nuclear magnetic resonance experiment, and the gas–water two-phase distribution state and displacement efficiency in coal with different porosity under different gas injection pressures are accurately characterized. The fractal dimension of liquid phase distribution under different gas injection pressures was obtained through experiments, and the gas–water two-phase migration law is inverted according to it. Finally, the gas–water two-phase migration mechanism inside the fractal structure of coal was obtained. The results are as follows: 1. Gas will first pass through the dominant pathway (the composition of the dominant pathway is affected by porosity) and it will continue to penetrate other pathways only when the gas injection pressure is high. When the gas injection pressure is low, the displacement occurs mainly in the percolation pores. With the increase in gas injection pressure, the focus of displacement gradually shifts to the adsorption pore. 2. As the gas injection pressure increases, the displacement efficiency growth rate is relatively uniform for the high-porosity coal samples, while the low-porosity coal samples show a trend of first fast and then slow growth rates. When the gas injection pressure reaches 7 MPa, the displacement efficiency of high-porosity coal samples exceeds that of low-porosity coal samples. 3. With the increase in gas injection pressure, the fractal dimension of the adsorption pore section and the seepage pore section shows an increasing trend, but the fractal dimension of the adsorption pore section changes faster, indicating that with the increase in gas injection pressure, gas–water two-phase displacement mainly occurs in the adsorption pore section.

Hydraulic fracturing of a roof can attenuate the strong mineral pressure and stress transfer by creating a series of parallel artificial fractures in it, and different forms of hydraulic parallel fractures have different effects on the attenuation effect of the coal rock system. In this regard, this study investigated the influence law of different forms of parallel pre-cracks on the mechanical properties of a rock–coal combined body through PFC^2D numerical simulation, and the following conclusions were obtained. Parallel pre-cracks significantly affected the mechanical properties of the rock–coal combined body. The stress–strain curve of the coal–rock assemblage containing Parallel pre-cracks changes significantly following reductions in peak strength, peak strain, elastic modulus, and crack initiation stress. The closer the angle θ between the single set of parallel pre-cracks and the horizontal is to 30°, the longer the length, L, and the lower the peak strength of the specimen, the peak strain, the modulus of elasticity, and the crack initiation stress. Macroscopic damage patterns are given for a rock–coal combined body containing single sets of parallel pre-cracks of different parameters, with coal sample damage, coal–rock sample damage and rock sample damage; a rock–coal combined body containing parallel pre-cracks is divided into three modes of fracture initiation when pressurized. These are the cracking of the coal sample, the simultaneous cracking of the parallel pre-crack tips in coal and rock samples, and the cracking of the parallel pre-crack tips in rock samples.

In order to study the local deformation of an anchor bolt and the improvement in the shear strength of a structural surface under the misalignment of an anchorage structure surface, FLAC^3D software was used to simulate granite, sandstone, and coal specimens with anchorage angles of 90° to analyze the damage of the anchoring agent and the changes in the local axial and shear forces of the anchor bolts with the misalignment of the structural surface. The results show that the anchor bolt near the structural surface had significant local characteristics with the misalignment of the structural surface; that is, the length of the local deformation area of the bolt was approximately equal to the length of the damaged area of the anchoring agent, and the stress on the anchor bolt was in a coupled tensile–shear stress state when the bolt reached the yield state. For the fully grouted bolts, it was this significant local feature that made the shear strength of the structural surface increase rapidly under a small shear displacement so that the structural surface reached a stable state. The improvement in the shear strength of the anchoring structural surface was caused by the misalignment of the structural surface. This is referred to as the passive improvement of the shear strength of the anchoring structural surface, which is the mechanism of the bonding section anchor to control the shear displacement of the structural surface and realize the stability of the rock mass.