【作者】 李斌；

【导师】 柏建彪；

【作者基本信息】 中国矿业大学 ， 采矿工程， 2022， 硕士

【摘要】 高瓦斯或煤与瓦斯突出煤层掘进速度小于回采速度的现状使矿井采掘关系紧张。大量工程实践表明,制约掘进速度的关键点主要有2个:（1）支护速度慢,掘支体系不匹配;（2）空顶距离小,掘支循环步距小、辅助时间长、导致掘进效率低。因此在保证施工安全的前提下,确定合理最大空顶距,研究掘进迎头顶板稳定控制机理、合理支护参数及其施工工序,对于提高巷道掘进效率,实现煤炭高产高效具有重要意义。本文基于李村煤矿3301进风底抽巷的工程背景,综合采用现场调研、室内试验、理论分析、数值计算及现场试验等研究方法,详细分析了底抽巷围岩地质力学特性、掘进迎头空顶区顶板力学结构及其稳定性的主要影响因素及其影响规律,揭示空顶区顶板变形破坏规律。研究确定了底抽巷的合理布置方案,并提出了高效掘支分步支护控制技术,并得到成功应用,取得以下主要成果:（1）根据底板滑移线场理论计算求得3301工作面底板破坏深度为9.73m。通过构建FLAC^3D数值模型,分析了煤层底板垂直应力分布特征及动压作用下底抽巷围岩变形破坏规律,基于岩石强度、掘进速度等,综合确定底抽巷布置于煤层下方10m、外错第一个工作面顺槽15m。（2）建立了掘进迎头空顶区顶板的薄板力学模型,推导了薄板模型的挠度、应力及掘进迎头极限空顶距表达式,并以顶板拉破坏准则为依据确定了掘进迎头的理论极限空顶距为2.97m。影响极限空顶距的灵敏度由大到小依次为空顶区浅部顶板发生离层的岩体厚度、空顶区顶板上覆载荷、巷道所在岩层强度。（3）分析了空顶距离、循环步距及支护强度等因素对空顶区顶板稳定性的影响规律:掘进迎头空顶距对空顶区巷道围岩稳定性影响较为显著,当空顶距大于4.0m后,空顶区顶板及帮部垂直应力均会出现减小的趋势及向深部转移,增大巷道围岩变形;空顶区顶板中部下沉量随着循环步距的增大呈线性增加,在空顶区后部顶板下沉量增幅随着循环步距的增大而增大;增加支护强度可有效改善顶板的承载状态,同时对支护区顶板有着明显的控制效果。（4）提出了提高掘进速度的锚杆及时支护、锚索滞后支护的高效掘支分步支护控制技术,确定了锚索滞后支护距离为10m、锚杆间排距为900×1400mm,锚索布置方式为2-1-2等关键支护参数。（5）确定了李村煤矿3301进风底抽巷采用2.8m大循环步距及锚索滞后10m的高效掘支方案;现场矿压观测及掘进效率实测表明,锚杆锚索有效协同承载、保障巷道安全稳定,掘进速度提高了40%以上。本论文中图56幅、表7个、参考文献86篇。更多还原

【Abstract】 The current situation of high gas or coal and gas outburst coal seam driving speed is less than the recovery speed,which makes the mine mining relationship tense.A large number of engineering practices show that there are two key points restricting the excavation speed:(1)the support speed is slow and the excavation branch system does not match;(2)the empty top distance is small,the excavation branch cycle step is small,the auxiliary time is long,and the excavation efficiency is low.Therefore,on the premise of ensuring construction safety,it is of great significance to determine the reasonable maximum ceiling distance,study the stability control mechanism of the head-head roof,reasonable support parameters and construction procedures,which are of great significance to improve the tunneling efficiency and achieve high coal production and high efficiency.Based on the engineering background of the 3301 air-inlet bottom-drawing roadway in Licun Coal Mine,this thesis comprehensively adopts research methods such as on-site investigation,laboratory test,theoretical analysis,numerical calculation and field test,etc.The mechanical structure of the roof in the roof area and the main influencing factors and influence laws of its stability reveal the deformation and failure laws of the roof in the empty roof area.The reasonable layout scheme of the bottom-drawing roadway was determined through research,and the high-efficiency branch excavation step-by-step support control technology was proposed,which was successfully applied,and the following main achievements were obtained:(1)According to the theoretical calculation of the slip line field of the floor,the failure depth of the floor of the 3301 working face is 9.73 m.By constructing the FLAC3 D numerical model,the vertical stress distribution characteristics of the coal seam floor and the deformation and failure law of the surrounding rock of the bottom-drawing roadway under the action of dynamic pressure are analyzed.A working face is 15 m along the groove.(2)The thin-plate mechanical model of the roof in the head-end void area is established,and the deflection,stress and the expression of the limit head-head distance of the thin-plate model are deduced.Based on the roof tensile failure criterion,the theoretical limit space of the head-head is determined.The top distance is 2.97 m.The sensitivity that affects the ultimate void-roof distance,from large to small,is the thickness of the rock mass that occurs in the shallow roof of the void-roof area,the overlying load of the roof in the void-roof area,and the strength of the roadway roof.(3)The influence law of the empty roof distance,cyclic step distance and support strength on the roof stability of the empty roof area is analyzed: the head-on empty head distance has a significant influence on the stability of the surrounding rock of the roadway in the empty roof area.After more than 4.0m,the vertical stress of the roof and the side of the empty roof area will decrease and transfer to the deep part,which will increase the deformation of the surrounding rock of the roadway.Linear increase,the increase of the subsidence of the roof at the rear of the empty roof area increases with the increase of the cycle step;increasing the support strength can effectively improve the bearing state of the roof,and at the same time has a significant control effect on the roof in the support area.(4)Propose the high-efficiency excavation branch support control technology of bolt timely support and anchor cable lag support to increase the driving speed,and determine the anchor cable lag support distance of 10 m and the row spacing between bolts to be 900× 1400 mm,the anchor cable arrangement is 2-1-2 and other key support parameters.(5)It is determined that the 3301 air-inlet and bottom-drawing roadway of Licun Coal Mine adopts a high-efficiency excavation plan with a large circulation step distance of 2.8m and an anchor cable lag of 10m;the on-site mine pressure observation and the actual measurement of the excavation efficiency show that the bolt and anchor cable effectively cooperate to carry and guarantee The roadway is safe and stable,and the driving speed is increased by more than 40%.In this thesis,there are 56 figures,7 tables,and 86 references.更多还原