错层位外错式沿空掘巷机理及相邻巷道的立体化联合支护技术

针对特厚煤层中应用错层位外错式沿空掘巷与相邻巷道的立体化联合支护机理展开研究。首先理论分析了错层位外错式沿空掘巷布置特点，发现:① 沿煤层顶板及起坡下方的三角煤体保持稳定且对实体煤提供侧向支承应力σx作用，因此一侧采空情况下实体煤从上至下出现了新的变化，即侧向支承应力从0增加至σx;② 依据巷道布置层位与侧向支承应力的不同，将特厚煤层从上至下进行了分区，分别为一侧采空实体煤分区、过渡区与弹性区;③ 布置接续工作面沿空掘巷时，其围岩处于弹性状态，巷道顶部属于一侧采空实体煤内的破碎区、塑性区，因此承载小，也即实现了围岩稳定与载荷低二者之间的统一。利用相邻两条巷道高、低不同这一立体化空间关系，提出相邻巷道的立体化联合支护技术，分析其特点包括:① 沿煤层顶板布置巷道，主动支护可打入深部稳定岩层内;② 通过加强一侧采空实体煤内的加固作用，为相邻沿煤层底板巷道提供锚固点;③ 底板沿空掘巷顶部支护体可深入顶板岩层、联合锚固区与过渡区，可实现全长锚固以更加充分发挥支护作用。为了验证前述理论成果，采用数值模拟进行计算分析，发现:① 采用错层位外错式沿空掘巷技术，显著改善了沿空巷道围岩性质与应力分布现状，实现了低应力与围岩稳定二者的统一，且验证了前述对错层位外错式沿空掘巷特厚煤层纵向分区的成果;② 零原岩应力场条件下联合支护效果较为明显，沿顶巷道顶板锚索深入基本顶段、巷帮一侧锚固段、沿底巷道顶板锚索深入岩层内与联合锚固区均出现应力集中作用，显著改善了沿底巷道单巷支护受顶煤厚度限制无法形成有效锚固点的现状;③ 在对实际工程背景的数值计算发现，与沿底巷道围岩大范围破坏且无法控制相比，采用错层位外错式沿空掘巷相邻巷道联合支护技术可显著控制沿空巷道围岩破坏范围。

Mechanism of external-misaligned stagger arrangement of roadway driven along next goaf and the 3D combined support technique of adjacent roadway

Based on the 3D(three-dimensional)combined support mechanism of external-misaligned stagger arrangement of roadway driven along next goaf and adjacent roadway in the ultra-thick coal seam,the roadway layout characteristics are theoretically analyzed and found that:① The roof of coal seam and triangular coal body could keep stable,providing the lateral abutment stress σx to solid coal.Therefore,the next goaf side of solid coal has been changed from top to bottom,that is,the lateral abutment stress increases from 0 to σx.② According to the difference in roadway layout and lateral abutment stress,the ultra-thick coal seam from top to bottom is divided into the next goaf side of solid coal zone,the transition zone,and the elastic zone.③ When arranging a continuous working face roadway driven along next goaf,its surrounding rock is in an elastic state,the roof of roadway carries small load,which locates on the crushing zone and the plastic zone in the next goaf side of the solid coal.The roadway achieves the unity of the stability of the surrounding rock and the low-load.Then the authors propose a 3D combined support technique of adjacent roadway by using high and low roadways of different 3D spatial relationship.Its characteristics of the proposed technique are as follows:① The active support body has been carried into the deep stable rock stratum in the roadway driven along the roof of the coal seam.② By strengthening the reinforcement of the next goaf side of solid coal zone,it provides an anchor point for the adjacent roadway along the bottom in the coal seam.③ The support body in the top of the roadway driven along next goaf has been carried into the deep roof of rock stratum by combining anchorage zone and transition zone,which can achieve a full-length anchorage to fully support function.In order to test the theoretical results,the numerical simulation analysis shows:① Through the use of external-misaligned stagger arrangement roadway driven along next goaf technique,it achieves the unity of stability of surrounding rock and low-stress,and tests the results of external-misaligned stagger arrangement roadway driven along next goaf’s vertical division.② Under the condition of obviously combined support effect of zero-rock stress field,the top of roadway driven along roof’s anchor cables has been carried into the deep basic roof and its roadway’s side of anchorage zone.The top of the roadway driven along next goaf’s anchor cables has been carried into the deep roof of rock stratum and combined anchorage zone,and stress concentration occurs in these zones.This technique forms an effective anchor point,which is not limited by the thickness of top coal.③ Through numerical calculation on a field engineering project,it is found that the proposed technique is capable of effectively controlling the large-scale destruction of the surrounding rock roadway than previously used ones.The results can provide a reference for ultra-thick coal seam mining.