In order to control the deformation of surrounding rock in deep high-stress and intense-mining roadways, taking a deep coal roadway with continuous deformation as an example, the characteristics of crustal stress, coal strength, and mining influence of roadway are obtained by underground tests. The combined failure mechanism of coal roadway surrounding-rock is revealed by differential stress of deep and shallow anchor cables. We propose that the improvement of surrounding rock control for coal roadway is adopting the coupling control technology of anchoring and unloading. The stress distribution and evolution laws of lateral surrounding rock of unloading holes are obtained by numerical simulation and theoretical calculation, and reasonable unloading-hole spacing of 4.0 m is comprehensively determined. A mechanical model of roadway roof beam under fixed support at both ends is constructed and the important role of anchor cable beam-truss in controlling the stability of coal roadway is obtained. The rationality of coupling control technology of anchoring and unloading and parameters has been verified by engineering test and mine pressure observation, providing technical references for surrounding rock control in deep intense-mining and large-deformation roadways.

The soft-rock roadways in kilometer-deep coal mines are often damaged by large deformation and have to be periodically expanded and repaired, which seriously restricts the safe and efficient production of coal mines. A typical soft-rock roadway in a kilometer-deep coal mine is selected as the engineering, and the main reasons for roadway deformation are analyzed, and the ground stress and mechanical characteristics are obtained. The Flac^3D numerical model, which can accurately reflect the deformation characteristics of surrounding rock in kilometer-deep soft-rock roadway, has been constructed, and the evolution laws of stress field and its damage mechanism have been analyzed with the vertical stress, vertical displacement and plastic zone. The damage range of roadway surrounding rock is elucidated by endoscopic observation, and the synergistic control technology of support and modification is proposed. The on-site engineering application is carried out in Qianyingzi Coal Mine, and the results show that the collaborative control technology of support and modification ensures the safety and stability of deep soft-rock roadway. The results provide important and preliminary technical references for realizing safe and intelligent mining in coal mines.

Fully grouted rock bolts are widely used in civil engineering and mining engineering, playing a significant role in keeping the stability and safety of excavations. In this paper, the load transfer mechanism of fully grouted rock bolts was studied with an analytical model. A trilinear model was used to depict the bond‐slip behaviour of the bolt/grout interface. The displacement of the confining medium was involved in this analytical model. Then, the shear stress propagation along the bolt/grout interface was analysed in the elastic, elastic‐softening, elastic‐softening‐debonding, softening‐debonding, and debonding stages. Experimental pull‐out tests were used to validate this analytical model. There was a good correlation between experimental and analytical results. A parametric study was conducted to evaluate the influence of Young’s modulus of the confining medium, the shear strength of the bolt/grout interface, and the residual shear strength of the bolt/grout interface on the load transfer performance of rock bolts. The results show that increasing Young’s modulus of the confining medium was beneficial for improving the load transfer performance of rock bolts. However, once Young’s modulus of the confining medium was beyond a critical limit, it had marginal effect on the peak load of rock bolts. Furthermore, increasing the shear strength of the bolt/grout interface and the residual shear strength of the bolt/grout interface led to rising of the peak load of rock bolts. However, compared with the residual shear strength of the bolt/grout interface, increasing the shear strength of the bolt/grout interface had more apparent effect in improving the peak load of rock bolts.

In response to the shortcomings of the small support scope of single anchor cable and failure to fully utilize the joint anchoring for surrounding rock, this study summarizes the classification of basic support forms for coal mine roadway surrounding rock and clarifies the main composition and mechanism of anchor cable beam‐truss structure. A mechanical model of roadway roof beam under the conditions of no support, single anchor cable support, and anchor cable beam‐truss structure support at both ends of the roadway is constructed, and the force and maximum bending moment of roadway roof beam under different support types are compared and analyzed. The study clarifies the bending moment reduction rate of anchor cable beam‐truss structure relative to unsupported and single anchor cable support and combines numerical simulation to analyze the response laws of anchor cable beam‐truss to prestress distribution and stress shell in roadway surrounding rock. Based on the on‐site engineering application of a typical large‐deformation mining roadway in a deep coal mine, the important role of anchor cable beam‐truss structure in ensuring the safety and stability of surrounding rock in deep high‐stress and intense‐mining large‐deformation roadway is revealed, providing technical references for surrounding rock control of similar conditions in deep roadways.

Aiming at the problem of surrounding rock deformation and failure of mining roadway and its control, a mechanical model of the circular roadway under the mining environment is established, and the implicit equation of the plastic zone boundary is derived. By analyzing the morphologic evolution law of the surrounding rock plastic zone in the mining roadway, the key factors affecting the morphologic change of the plastic zone are obtained, that is, the magnitude and direction of principal stress. The influence law of the magnitude and direction of principal stress on the plastic zone of the mining roadway is analyzed by using numerical simulation software, and the deformation and failure mechanism of surrounding rock of the mining roadway is revealed. The results showed that the size and morphology of the plastic zone were closely related to the confining pressure ratio (η). Taking the boundary of η valuing 1, the larger or smaller η value was, the more serious the deformation and failure of surrounding rock would be; the morphology of the plastic zone changed with the deflection of the principal stress, with the location of the maximum plastic zone influenced by the principal stress direction. For the surrounding rock control in the mining‐influenced roadway, it is advised to take the following methods: firstly, it is necessary to consider how to reduce or remove the influence of mining on surrounding rock, improve the stress environment of surrounding rock, and reduce the failure depth of the plastic zone, so as to better maintain the roadway. Secondly, in view of the deformation and failure characteristics of the mining roadway, the fractional support method of “yielding first and then resisting” should be adopted, which applies the cable supplement support after mining instead of the one‐off high‐strength support during roadway excavation, so as to control the malignant expansion of the surrounding rock plastic zone and prevent roof falling accidents.

Aiming to address the challenges of determining the coal pillar’s width and managing the significant deformation of the surrounding rock in the deep gob-side entry driving, the limiting equilibrium zone theory, employing the operational area of Dongpang Mine 21110 as the engineering setting, states that a coal pillar’s appropriate width in the gob-side entry driving falls between 7.9 and 9.8 m. The pattern of vertical stress distribution and the extent of the plastic zone in the roadway for coal pillar widths of 7.0 m, 8.0 m, 9.0 m, and 10.0 m are analyzed, respectively, investigated using the numerical simulation method of FLAC^3D. The acceptable coal pillar width in the deep gob-side entry driving is 8.0 m. Combined with the roadway surrounding rock borehole inspection results, the fracture development condition of the roadway’s full-face surrounding rock is determined, and the asymmetric aberration characteristics, with significant surrounding rock damage depth at the coal pillar flank location, are obtained. Based on the theoretical calculations, an integrated proposal for a “non-symmetrical bolt and cable anchor” coupling support scheme for the surrounding rock in the gob-side entry driving is put forward. This was applied at the Dongpang coal mine site. Engineering practice shows that leaving an 8.0 m coal pillar width and adopting the “non-symmetrical bolt and cable anchor” support system design can control the deformation of the surrounding rock in the track entry at a reasonable range, which ensures the stability of the surrounding rock in the gob-side entry driving.