This paper studies the width of narrow coal pillars, mining-induced failure characteristics, and surrounding rock control effect of gob-side entry driving (GED) adjacent to 2-1208 filling working face with an approximately 900 m depth. Laboratory experiments, numerical simulations, loosening circle tests, and engineering practices are conducted. The mechanical properties of the filling body, the distribution and evolution law of the second invariant deviatoric stress (J 2 ), and the variation in the plastic zone of the surrounding rock in GED are studied. The conditions of various coal pillar widths and the gob backfilled or not of the adjacent working face are also considered.

Regarding the issue of intense mining pressure appearing in the underlying gateway below the remaining coal pillar in the close‐distance coal seam (the remaining coal pillar is perpendicular to the underlying section coal pillar), 401 working face is used as the engineering background. Field measurements, laboratory experiments, numerical simulations, and engineering verification techniques are used to study the abutment pressure's evolution properties and the plastic zone's propagation laws before and after the underlying coal seam roadway experienced the mining impact. The conclusions are as follows: ① The maximum plastic area on the two sides and the roof of the roadway underlying the gob are up to 2 and 1.5 m, whereas the maximum plastic area on the two sides and the roof of the roadway underlying the remaining coal pillar are up to 5 and 4.5 m, respectively. Moreover, the plastic area extends along the two sides, and the section coal pillar is completely broken when the working face is mined below the remaining coal pillar. ② The stress increase coefficient K in the overlap area of the remaining coal pillar and the underlying section coal pillar reaches 3.4 when the mining face penetrates the underlying remaining coal pillar and the advance abutment pressure is overlaid with the concentrated stress of the coal pillar. ③ When the underlying working face is mined to 4, −2, −8, and −14 m away from the remaining coal pillar, the damage range of the roadway 5–10 m ahead increases in turn. At the same time, the maximum plastic area of the roof passes through the plastic area of the upper coal seam floor. Therefore, the underlying and transition areas on both sides of the remaining coal pillar are divided into Area I (15 m) → Area II (the most complicated area to control under the remaining coal pillar, 20 m) → Area III (25 m) according to the width. Furthermore, the divisional differentiated combined control technology of channel steel truss anchor cable with joint double‐way locking control function of roof and coal pillar in Areas I and III, while channel steel truss anchor cable with joint double‐way locking control function of roof and side + high resistance integral door‐type support is proposed in Area II. Field engineering practice shows that the deformation of the roadway surrounding rock can be controlled within 210 mm after adopting the above divisional combined control technology. Finally, the mining operation can safely and efficiently pass through the remaining coal pillar. The research results have important reference values for surrounding rock control of mining roadways in the overlapping area of similar “+”‐type cross‐working face.

In deep underground mining, achieving stable support for roadways along with long service life is critical and the complex geological environment at such depths frequently presents a major challenge. Owing to the coupling action of multiple factors such as deep high stress, adjacent faults, cross-layer design, weak lithology, broken surrounding rock, variable cross-sections, wide sections up to 9.9 m, and clusters of nearby chambers, there was severe deformation and breakdown in the No. 10 intersection of the roadway of large-scale variable cross-section at the − 760 m level in a coal mine. As there are insufficient examples in engineering methods pertaining to the geological environment described above, the numerical calculation model was oversimplified and support theory underdeveloped; therefore, it is imperative to develop an effective support system for the stability and sustenance of deep roadways. In this study, a quantitative analysis of the geological environment of the roadway through field observations, borehole-scoping, and ground stress testing is carried out to establish the FLAC 3D variable cross-section crossing roadway model. This model is combined with the strain softening constitutive (surrounding rock) and Mohr–Coulomb constitutive (other deep rock formations) models to construct a compression arch mechanical model for deep soft rock, based on the quadratic parabolic Mohr criterion. An integrated control technology of bolting and grouting that is mainly composed of a high-strength hollow grouting cable bolt equipped with modified cement grouting materials and a high-elongation cable bolt is developed by analyzing the strengthening properties of the surrounding rock before and after bolting, based on the Heok-Brown criterion. As a result of on-site practice, the following conclusions are drawn: (1) The plastic zone of the roof of the cross roadway is approximately 6 m deep in this environment, the tectonic stress is nearly 30 MPa, and the surrounding rock is severely fractured. (2) The deformation of the roadway progressively increases from small to large cross-sections, almost doubling at the largest cross-section. The plastic zone is concentrated at the top plate and shoulder and decreases progressively from the two sides to the bottom corner. The range of stress concentration at the sides of the intersection roadway close to the passageway is wider and higher. (3) The 7 m-thick reinforced compression arch constructed under the strengthening support scheme has a bearing capacity enhanced by 1.8 to 2.3 times and increase in thickness of the bearing structure by 1.76 times as compared to the original scheme. (4) The increase in the mechanical parameters c and φ of the surrounding rock after anchoring causes a significant increase in σt; the pulling force of the cable bolt beneath the new grouting material is more than twice that of ordinary cement grout, and according to the test, the supporting stress field shows that the 7.24 m surrounding rock is compacted and strengthened in addition to providing a strong foundation for the bolt (cable). On-site monitoring shows that the 60-days convergence is less than 30 mm, indicating that the stability control of the roadway is successful.

Close-distance coal seams are widely distributed in China, and the mining of overlying coal seams leads to floor damage. To grasp the properties and the fracture spans of the damaged main roof in the underlying coal seam, combining the calculation of the floor damage depth with rock damage theory and the formulas for calculating the first and periodic weighting intervals of the damaged main roof and the instability conditions of the damaged key blocks are obtained. Three interaction stability mechanics models are proposed for key blocks with different properties of the upper and lower main roof, and the instability conditions of the lower damaged key blocks are obtained when the fracture lines overlap. When combined with a specific example, the field monitoring verified the calculation results. The research results are as follows: (1) The first and periodic weighting intervals, horizontal thrust between blocks, and critical load of instability of the damaged main roof are significantly reduced. Still, there are differences in its reduction under different loads, rotation angles, and lumpiness. (2) When the fracture lines of the upper and lower main roofs overlap, the stability of the damaged key blocks is the lowest. There are three linkage stability regions in the critical load curves of the two key blocks. (3) In this case, the damage equivalent of the main roof is 0.397, which belongs to the local damage type. Its first and periodic weighting intervals are 40 m and 16 m, which is 22% and 24% less than when there is no damage. (4) A supporting load of 0.489 MPa is required to maintain the stability of the upper key block, and the lower damaged key block is prone to rotary and sliding instability during the first and periodic weighting, respectively. Thus, the supports need to bear a total of 0.988 MPa and 0.761 MPa to maintain the stability of the two key blocks simultaneously. The ground pressure data monitored on-site is in accord with the calculation results.

In multi-seam mining, the residual coal pillar (RCP) in the upper gob has an important influence on the layout of the roadway in the lower coal seam. At present, few papers have studied the characteristics of the surrounding rock of gob-side entry driving (GED) with different coal pillar widths under the influence of RCP. This research contributes to improving the recovery rate of the extra-thick coal seam under this condition. The main research contents were as follows: (1) The mechanical parameters of the rock and coal mass were obtained using laboratory experiments coupled with Roclab software. These parameters were substituted into the established main roof structure mechanics model to derive the breakage position of the main roof with the influence of RCP, and the rationality of the calculation results was verified by borehole-scoping. (2) Based on numerical simulation, the evolution laws of the lateral abutment stress in the lower working face at different relative distances to the RCP were studied. FLAC3D was used to study the whole space-time evolution law of deviatoric stress and plastic zone of GED during driving and retreating periods with various coal pillar widths under the influence of RCP. (3) The plasticization factor P was introduced to quantify the evolution of the plastic zone in different subdivisions of the roadway surrounding rock, so as to better evaluate the bearing performance of the surrounding rock, which enabled a more effective determination of the reasonable coal pillar width. The field application results showed that it was feasible to set up the gob-side entry with an 8 m coal pillar below the RCP. The targeted support techniques with an 8 m coal pillar could effectively control the surrounding rock deformation.