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.

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.

Close-distance coal seams are widely distributed over China, and the coal pillars left by the overlying coal seams affect the retracement channel of the underlying coal seam in the stopping stage. Based on the engineering background of close-distance seam mining in a coal mine, the reasonable position of the underlying coal seam's stopping line and the support method of the large section roadway during stopping are investigated using field measurements, similar simulation experiments, and numerical simulations. There are three types of location relationships between the stopping line of the underlying coal seam and the stopping line of the overlying coal seam: "externally staggered with the upper stopping line" (ESUL, stops mining under the overlying goaf), "overlapped with upper stopping line" (OUL), and "internally staggered with the upper stopping line" (ISUL, ISUL-SD for shorter internal staggered distances, ISUL-LD for longer ones). There are different stress arch structures in the overlying strata of the above three positions, and the stress arch evolution process exists in the process of ESUL → OUL → ISUL-SD → ISUL-LD: a front and rear double stress arch structure → the front arch gradually decreases → the front arch dies out, and the double arch synthesizes the single arch → the single-arch range expands → the nested double arch. The relationship between the stress arch structure and the position of the stopping line is evaluated as follows: (1) ESUL: the stress concentration in the roof plate of the retracement channel of the underlying coal seam is the highest, because the overburden block of the extensive collapse zone acts directly on the roof plate of the retracement channel, resulting in relative difficulties in roof support. (2) OUL: although the retracement channel roof pressure is minimal, the overlying rock structure has the potential for rotation or slippage instability. (3) ISUL-SD: the pressure on the roof of the retracement channel is small and the overburden structure is stable, which is conducive to the safe retraction of the support and not limited by the width of the end-mining coal pillar. (4) ISUL-LD: it is basically the same as the condition of stopping under the non-goaf; however, it has a limitation on the width of the end-mining coal pillar. The location of the stopping line is selected as ISUL-SD, and the retraction process of the self-excavating retraction channel was adopted. A partition asymmetric support scheme which is proven by field practice is proposed, through a comprehensive analysis of the pre-stress field simulation of the support scheme, based on the different control requirements of the roof above the support and the roof of the retracement channel in the stopping area. This method realizes safe and smooth withdrawal of the support.

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.