为解决某矿在高应力软岩条件下巷道变形严重、难支护等问题，采用理论计算、数值模拟和现场实测等方法对巷道布置方式和支护方式进行优化，采用错层位沿空掘巷巷道布置方式，建立3个铰接岩块间力学模型，确定了作用在煤柱上的应力及应力集中系数，进而建立了缓斜中厚煤层采空区实体煤侧力学模型，通过理论推导、计算得出极限平衡区宽度，确定了沿空掘巷巷道位置，并提出了“相邻巷道联合支护技术和非对称控制技术”相结合的综合围岩控制技术。研究结果表明：作用在煤柱上的应力为55.71 MPa，应力集中系数为4.60，进而得出煤层垂向应力集中系数分量为4.33、倾向应力集中系数分量为1.57，极限平衡区宽度为11.35 m，确定煤柱宽度为5 m，数值模拟显示采用错层位综合围岩控制技术，1218回风巷顶板垂直应力大于原岩应力，可承载一部分上覆岩层质量，而传统巷道支护垂直应力低于原岩应力，说明顶板已较为破碎，不具备承载能力，传统支护右帮低于原岩应力，围岩破碎。2种支护方式煤柱均发生破坏，采用综合围岩控制技术时顶板有一部分未发生破坏，右帮破坏范围小，采用传统支护时顶板破坏范围较大，右帮破坏大，说明采用综合围岩控制技术可提高围岩强度和抗破坏能力。通过现场监测，在巷道掘进75 d后围岩变形量减缓。顶板最大下沉量为74 mm，底鼓量最大为32 mm，煤柱帮侧位移最大为58 mm，实体煤帮侧最大位移为45 mm，说明采用综合围岩控制技术可有效控制围岩变形量。

To solve the problem in a mine under the condition of high stress soft rock roadway deformation is serious，the problem such as difficult to support，the theoretical calculation，numerical simulation and spot observation and other methods to optimize roadway layouts and supporting way，using the wrong layer of gob-side entry driving arrangement，three hinged between the rock mechanics model is set up，to determine the effect on the coal pillar stress and stress concentration coefficient，and then slowly inclined thick coal seam in the goaf is established the entity coal side mechanical model，through theoretical derivation，calculation of limit equilibrium zone width，determine the gob-side entry driving position,The comprehensive surrounding rock control technology which combines “adjacent roadway joint support technology and asymmetric control technology” is put forward.The results show that:role stress on coal pillar is 55.71 MPa，the stress concentration factor is 4.60，and it is concluded that coal seam vertical stress concentration coefficient of components is 4.33,tendency of stress concentration coefficient is 1.57，the limit equilibrium section width of 11.35 m，determine the coal pillar width is 5 m，numerical simulations show the wrong layer integrated control technology of surrounding rock，1 218 vertical stress is greater than the original rock stress，return air roadway roof part carry weight of overburden，and traditional roadway support vertical stress is lower than the original rock stress，roof has been broken，do not have bearing capacity，the traditional support help right below the original rock stress,The surrounding rock is broken.Both of the two supporting methods of coal pillars are damaged，and the comprehensive surrounding rock control technology is adopted.Part of the roof is not damaged，while the right side is small in the scope of damage，while the traditional supporting roof is large in the scope of damage，and the right side is large in the scope of damage，which indicates that the comprehensive surrounding rock control technology can improve the strength and anti-damage ability of the surrounding rock.Through on-site monitoring，the deformation of surrounding rock was reduced after 75 days of roadway excavation.The maximum roof subsidence is 74 mm，the maximum floor heave is 32 mm，the maximum coal pillar side displacement is 58 mm，and the maximum solid coal side displacement is 45 mm，which indicates that the comprehensive surrounding rock control technology can effectively control surrounding rock deformation.

0 引言

1 工程概况

2 煤柱应力集中系数计算

   2.1 关键层确定

   2.2 基本顶周期及侧向垮落步距

   2.3 块体力学分析

3 窄煤柱宽度理论计算

   3.1 理论推导

   3.2 确定煤柱宽度

4 错层位沿空掘巷围岩控制技术

   4.1 围岩控制技术

   4.2 支护参数

5 数值模拟

   5.1 垂直应力

   5.2 塑性区分布分析

6 工业试验

7 结论