王家岭矿综放大断面剧烈采动影响煤巷强化控制研究

发布时间：2018-05-07 16:35

本文选题：剧烈采动煤巷 + 采动影响分区　；参考：《中国矿业大学(北京)》2017年博士论文

【摘要】：本文在现场矿压显现观测的基础上,通过岩石力学参数测定、力学建模、数值计算分析和工业性试验验证等方法研究大断面剧烈采动影响煤巷顶板破坏机理和围岩分区控制两个关键问题,分别对剧烈采动煤巷破坏特征、采动影响程度分区、采动影响煤巷上覆岩层结构的破坏及运移规律、巷道顶板离层错动机理及关键影响因素、高预应力桁架锚索双向控制机理、剧烈采动影响煤巷围岩分区段强化控制技术、王家岭典型大断面煤巷现场工业性试验等问题开展了一系列研究,主要研究成果如下:(1)基于实验室岩石物理力学特性试验结果,判定典型剧烈采动影响煤巷围岩等级为V类围岩,其自稳时间短,稳定性较差。现场调研发现:20102区段回风平巷440-550m区段受剧烈采动影响,巷道围岩严重失稳,0-440m区段暂未经受相邻工作面采动影响,但部分位置也发生较大破坏,如不加以控制,破坏程度会进一步加重。(2)综合巷道矿压显现调研和数值建模分析发现:综放面剧烈采动影响区域由工作面前方80m至后方130m范围;基于煤巷不同区段变形破坏特征,提出采动影响剧烈程度分区概念,并据此将20102煤巷划分为I-回采动压作用区、II-采掘联合动压作用区和III-采后静压作用区3个区段。(3)分析偏应力不变量综合研究指标得出:拉应变滞后转化形式是影响煤巷围岩采动影响剧烈程度的主导因素;相对于I-回采动压作用区和III-采后静压作用区,拉应变在II-采掘联合动压作用区滞后转化更为明显:煤巷顶板浅部围岩破坏较为严重,应变类型为拉应变;中位岩层范围内围岩保持较高程度畸变能,应变类型由拉应变向平面应变类型转化,表现出明显的滞后转化;深部围岩应变为单一压应变。(4)基于煤巷上覆岩层断裂后铰接特征,建立上覆岩块铰接结构力学模型,提出采动影响下巷道围岩所承受载荷的计算方法,q4为采动影响条件下煤柱帮所承受的均布载荷,Fs为采动影响时实体煤帮所承受的荷载,表达式为:(?)(5)运用离散元数值模拟软件,数值建模分析煤巷上覆岩层大结构运动规律,发现结构中岩块B的剧烈回转下沉运动将直接导致煤巷顶板在垂直方向急剧下沉,在水平方向促使基本顶向巷道内侧挤压剪切破坏,加大煤柱塑性区深度;煤巷围岩的破坏加剧覆岩结构的运动,甚至产生再次破断,更深的损坏了围岩的整体性;静压下煤巷围岩主要以垂直位移为主,剧烈采动影响下巷道围岩垂直位移进一步加剧,并伴有水平位移,垂直位移依然是围岩变形的主要形式。(6)依据材料力学中关于梁在垂直方向的挠度和水平方向的位移计算方法,发现煤巷顶板岩层错动产生于离层之后,只有当上位岩层的最大挠度小于下位岩层的最大挠度时才会产生岩层面的分离,综合剧烈采动影响煤巷覆岩结构特征,引入错动失稳系数KC,提出煤巷顶板煤岩层产生离层错动的判据:(?)(7)基于正交试验原理,设计离层错动影响因素数值模拟试验方案,分析数值计算结果得到影响煤层顶板巷道离层变形3个关键影响因素为:锚杆长度、顶煤强度和采掘关系;影响顶板水平错动的3个关键因素为:锚索角度、顶煤厚度和采掘关系;并详述出顶板离层错动与各因素之间的互馈关系;(8)针对剧烈采动影响煤巷离层错动破坏现象研发出新型高预应力桁架锚索结构,提出采用高预应力桁架锚索控制系统进行巷道围岩控制的改进方向,能够针对煤巷顶板垂直离层和水平错动变形做出双向的积极响应,尤其对大断面、采动影响剧烈、顶板软弱、高应力及悬顶面积大等复杂环境下的巷道围岩控制效果突出。(9)依据桁架锚索在工作状态下与顶板岩层紧密贴合的特点,建立顶板-桁架锚索组合梁结构力学模型,分别分析桁架锚索在垂直方向和水平方向对顶板煤岩层垂直挠度和水平错动的控制机理;建立桁架锚索结构承载模型,分析求解出桁架锚索材料的最低抗拉强度和安装时的预紧力,为支护方案设计奠定理论基础。(10)基于对煤巷采动影响剧烈程度分区的研究,提出了剧烈采动煤巷分区强化控制概念;分析煤巷各区段围岩的破坏特征,指出煤巷各区段相应的控制机理,运用数值建模分析方法与现场实践经验相结合,综合制定各区段支护方案;煤巷回采动压影响区:原有支护整修形成支护系统→采用桁架锚索控制顶板离层错动→补打锚索阻止塑性破坏扩展→单体柱提高剧烈采动安全储备;煤巷掘采联合动压作用区:提出了以双向控制为核心的“多支护结构”重建锚固体的修复系统;煤巷采后静压影响区:桁架锚索、单体锚索和锚杆相结合,多重支护结构相辅,预防本工作面回采时动压影响。(11)对各区段支护方案进行现场工业性试验,并对采动煤巷围岩控制效果进行顶板离层、表面位移监测,观测结果表明煤巷修复后,在相邻工作面剧烈采动影响下顶底板最大移近量约138mm,两帮最大移近量约为117mm,取得了良好的控制效果。
[Abstract]:In this paper, on the basis of the field observation of rock pressure, two key problems, such as rock mechanics parameter measurement, mechanical modeling, numerical analysis and industrial test verification, are studied. The failure mechanism of coal roadway roof and the zoning control of surrounding rock are studied, and the damage characteristics of coal roadway and the influence degree of mining are divided respectively. In the area, the damage and movement of the overlying strata in the coal roadway, the mechanism of the roof separation and the key influencing factors, the two-way control mechanism of the high prestress truss anchor cable, the intensive mining influence on the strengthening control technology of the partition section of the coal roadway, and the industrial test of the coal roadway in Wangjialing's typical large section coal roadway have been carried out in a series of research. The main research results are as follows: (1) based on the experimental results of laboratory rock physical and mechanical properties, it is found that the grade of surrounding rock of coal roadway of typical drastic mining influence is V class rock, its self stabilization time is short and the stability is poor. It is found that the 440-550m section of the return air lane in the 20102 section is affected by severe mining, the roadway surrounding rock is seriously unstable, and the 0-440m section is temporary. It is not affected by the mining action of adjacent working face, but the partial position also has great damage, if not controlled, the damage degree will be further aggravated. (2) the comprehensive roadway mining pressure investigation and numerical modeling analysis found that the severe mining area of fully mechanized caving face is from the front side 80m to the rear 130m range; based on the different sections of the coal roadway deformation and destruction special According to the concept of zoning, the 20102 coal lanes are divided into I- mining dynamic pressure area, II- mining joint dynamic pressure area and III- postharvest static pressure area. (3) analysis of the comprehensive study index of partial stress invariants: the lag transformation of tensile strain is the influence of the drastic degree of coal roadway surrounding rock mining. The leading factors, compared with the I- mining dynamic pressure area and the III- postharvest static pressure area, the lag transformation of the tensile strain in the II- mining joint dynamic pressure area is more obvious: the surrounding rock in the shallow part of the coal roadway roof is more serious, the strain type is the tensile strain, the surrounding rock in the middle rock stratum keeps a high degree of distortion energy, and the strain type is from the tensile strain to the plane. The transformation of strain type shows obvious lag transformation, and the strain of deep surrounding rock is a single compressive strain. (4) based on the hinge joint characteristics of the overlying strata after the coal roadway, the mechanical model of the hinged structure of the overlying rock is set up, and the calculation method of the load subjected to the surrounding rock under the influence of the mining is put forward, and Q4 is the uniform load of the coal pillar under the mining influence conditions. Load, Fs is the load which the solid coal support is subjected to during the mining effect. The expression is: (5) using the discrete element numerical simulation software, the numerical modeling is used to analyze the large structure movement law of the overlying strata in the coal roadway. It is found that the violent turning and sinking movement of the rock block B in the structure will directly cause the roof of the coal roadway to sink sharply in the vertical direction and promote the basic in the horizontal direction. The collapse of the top to the inner side of the roadway increases the depth of the plastic zone of the coal pillar, and the destruction of the surrounding rock intensifies the movement of the overlying rock structure, and even breaks down again, and further damages the integrity of the surrounding rock; the surrounding rock of the coal roadway under the static pressure is mainly vertical displacement, and the vertical displacement of the surrounding rock is further aggravated under the influence of severe mining, accompanied by water. Horizontal displacement, vertical displacement is still the main form of deformation of surrounding rock. (6) according to the displacement calculation method of the deflection and horizontal direction of beam in the vertical direction of material mechanics, it is found that the fault movement of rock strata in the roof of the coal roadway is produced after the separation, and only when the maximum deflection of the upper strata is smaller than the maximum deflection of the lower rock layer, the rock layer will be produced. The characteristics of the overlying rock structure of coal roadway are influenced by the comprehensive drastic mining, and the misdynamic instability coefficient KC is introduced. The criterion of separation of coal rock strata from coal roadway is proposed. (7) based on the principle of orthogonal test, the numerical simulation test scheme is designed for the influence factors of the separation of the strata, and the results of numerical calculation are analyzed to get 3 deformations of the coal seam roof. The key influencing factors are the length of bolt, the strength of top coal and the relation of mining, and the 3 key factors that affect the horizontal dislocation of the roof are the angle of anchor cable, the thickness of top coal and the relation of mining, and the reciprocal feed relationship between the separation and the factors of the top plate, and (8) a new type of high prestress is developed for the strenuous mining influence of the separation of the coal lanes. The structure of truss anchorage cable is improved by using the high prestress truss cable control system to control the surrounding rock of the roadway. It can make a two-way positive response to the vertical separation and horizontal dislocation deformation of the roof of the coal roadway, especially the roadway under the complex environment such as the large section, the mining action, the roof is weak, the high stress and the overhanging area are large. The control effect of the surrounding rock is outstanding. (9) according to the characteristics of close fitting of the truss cable with the roof rock layer under the working condition, the structural mechanics model of the roof truss cable composite beam is established, and the control mechanism of the vertical deflection and horizontal dislocation of the roof coal rock in the vertical direction and the horizontal direction is analyzed respectively, and the structure bearing of the truss anchor cable is established. The model, analysis and solution of the minimum tensile strength of truss anchor cable materials and the pre tightening force at installation, laid a theoretical foundation for the design of support scheme. (10) based on the study of the zoning of the intensity of coal roadway mining, the concept of zoning strengthening control for strenuous mining coal roadway was put forward, and the damage characteristics of surrounding rock in each section of coal roadway were analyzed, and the various areas of coal roadway were pointed out. The corresponding control mechanism, combining the numerical modeling analysis method with the field practice experience, comprehensively formulates the support scheme of each section, and the impact zone of the coal roadway recovery dynamic pressure: the original support is refurbished to form the supporting system. Safety reserve, coal roadway mining and mining joint dynamic pressure area: a "multi support structure" reconstruction system is put forward with two-way control as the core, and the influence area of post mining stope static pressure: truss anchorage, combination of single anchor cable and anchor, multiple support structure supplemented to prevent the dynamic pressure of this working face. (11) support for each section The case carries on the field industrial test, and carries on the roof separation and the surface displacement monitoring to the control effect of the surrounding rock of the coal roadway. The observation results show that the maximum displacement of the top floor of the top floor is about 138mm and the maximum displacement of the two gang is about 117mm after the coal roadway is repaired, and the good control effect has been obtained.

【学位授予单位】：中国矿业大学(北京)
【学位级别】：博士
【学位授予年份】：2017
【分类号】：TD322;TD353

【参考文献】