發(fā)布時(shí)間：2018-05-07 16:35

  本文選題：劇烈采動(dòng)煤巷 + 采動(dòng)影響分區(qū)��； 參考：《中國礦業(yè)大學(xué)(北京)》2017年博士論文

【摘要】：本文在現(xiàn)場礦壓顯現(xiàn)觀測的基礎(chǔ)上,通過巖石力學(xué)參數(shù)測定、力學(xué)建模、數(shù)值計(jì)算分析和工業(yè)性試驗(yàn)驗(yàn)證等方法研究大斷面劇烈采動(dòng)影響煤巷頂板破壞機(jī)理和圍巖分區(qū)控制兩個(gè)關(guān)鍵問題,分別對劇烈采動(dòng)煤巷破壞特征、采動(dòng)影響程度分區(qū)、采動(dòng)影響煤巷上覆巖層結(jié)構(gòu)的破壞及運(yùn)移規(guī)律、巷道頂板離層錯(cuò)動(dòng)機(jī)理及關(guān)鍵影響因素、高預(yù)應(yīng)力桁架錨索雙向控制機(jī)理、劇烈采動(dòng)影響煤巷圍巖分區(qū)段強(qiáng)化控制技術(shù)、王家?guī)X典型大斷面煤巷現(xiàn)場工業(yè)性試驗(yàn)等問題開展了一系列研究,主要研究成果如下:(1)基于實(shí)驗(yàn)室?guī)r石物理力學(xué)特性試驗(yàn)結(jié)果,判定典型劇烈采動(dòng)影響煤巷圍巖等級為V類圍巖,其自穩(wěn)時(shí)間短,穩(wěn)定性較差。現(xiàn)場調(diào)研發(fā)現(xiàn):20102區(qū)段回風(fēng)平巷440-550m區(qū)段受劇烈采動(dòng)影響,巷道圍巖嚴(yán)重失穩(wěn),0-440m區(qū)段暫未經(jīng)受相鄰工作面采動(dòng)影響,但部分位置也發(fā)生較大破壞,如不加以控制,破壞程度會(huì)進(jìn)一步加重。(2)綜合巷道礦壓顯現(xiàn)調(diào)研和數(shù)值建模分析發(fā)現(xiàn):綜放面劇烈采動(dòng)影響區(qū)域由工作面前方80m至后方130m范圍;基于煤巷不同區(qū)段變形破壞特征,提出采動(dòng)影響劇烈程度分區(qū)概念,并據(jù)此將20102煤巷劃分為I-回采動(dòng)壓作用區(qū)、II-采掘聯(lián)合動(dòng)壓作用區(qū)和III-采后靜壓作用區(qū)3個(gè)區(qū)段。(3)分析偏應(yīng)力不變量綜合研究指標(biāo)得出:拉應(yīng)變滯后轉(zhuǎn)化形式是影響煤巷圍巖采動(dòng)影響劇烈程度的主導(dǎo)因素;相對于I-回采動(dòng)壓作用區(qū)和III-采后靜壓作用區(qū),拉應(yīng)變在II-采掘聯(lián)合動(dòng)壓作用區(qū)滯后轉(zhuǎn)化更為明顯:煤巷頂板淺部圍巖破壞較為嚴(yán)重,應(yīng)變類型為拉應(yīng)變;中位巖層范圍內(nèi)圍巖保持較高程度畸變能,應(yīng)變類型由拉應(yīng)變向平面應(yīng)變類型轉(zhuǎn)化,表現(xiàn)出明顯的滯后轉(zhuǎn)化;深部圍巖應(yīng)變?yōu)閱我粔簯?yīng)變。(4)基于煤巷上覆巖層斷裂后鉸接特征,建立上覆巖塊鉸接結(jié)構(gòu)力學(xué)模型,提出采動(dòng)影響下巷道圍巖所承受載荷的計(jì)算方法,q4為采動(dòng)影響條件下煤柱幫所承受的均布載荷,Fs為采動(dòng)影響時(shí)實(shí)體煤幫所承受的荷載,表達(dá)式為:(?)(5)運(yùn)用離散元數(shù)值模擬軟件,數(shù)值建模分析煤巷上覆巖層大結(jié)構(gòu)運(yùn)動(dòng)規(guī)律,發(fā)現(xiàn)結(jié)構(gòu)中巖塊B的劇烈回轉(zhuǎn)下沉運(yùn)動(dòng)將直接導(dǎo)致煤巷頂板在垂直方向急劇下沉,在水平方向促使基本頂向巷道內(nèi)側(cè)擠壓剪切破壞,加大煤柱塑性區(qū)深度;煤巷圍巖的破壞加劇覆巖結(jié)構(gòu)的運(yùn)動(dòng),甚至產(chǎn)生再次破斷,更深的損壞了圍巖的整體性;靜壓下煤巷圍巖主要以垂直位移為主,劇烈采動(dòng)影響下巷道圍巖垂直位移進(jìn)一步加劇,并伴有水平位移,垂直位移依然是圍巖變形的主要形式。(6)依據(jù)材料力學(xué)中關(guān)于梁在垂直方向的撓度和水平方向的位移計(jì)算方法,發(fā)現(xiàn)煤巷頂板巖層錯(cuò)動(dòng)產(chǎn)生于離層之后,只有當(dāng)上位巖層的最大撓度小于下位巖層的最大撓度時(shí)才會(huì)產(chǎn)生巖層面的分離,綜合劇烈采動(dòng)影響煤巷覆巖結(jié)構(gòu)特征,引入錯(cuò)動(dòng)失穩(wěn)系數(shù)KC,提出煤巷頂板煤巖層產(chǎn)生離層錯(cuò)動(dòng)的判據(jù):(?)(7)基于正交試驗(yàn)原理,設(shè)計(jì)離層錯(cuò)動(dòng)影響因素?cái)?shù)值模擬試驗(yàn)方案,分析數(shù)值計(jì)算結(jié)果得到影響煤層頂板巷道離層變形3個(gè)關(guān)鍵影響因素為:錨桿長度、頂煤強(qiáng)度和采掘關(guān)系;影響頂板水平錯(cuò)動(dòng)的3個(gè)關(guān)鍵因素為:錨索角度、頂煤厚度和采掘關(guān)系;并詳述出頂板離層錯(cuò)動(dòng)與各因素之間的互饋關(guān)系;(8)針對劇烈采動(dòng)影響煤巷離層錯(cuò)動(dòng)破壞現(xiàn)象研發(fā)出新型高預(yù)應(yīng)力桁架錨索結(jié)構(gòu),提出采用高預(yù)應(yīng)力桁架錨索控制系統(tǒng)進(jìn)行巷道圍巖控制的改進(jìn)方向,能夠針對煤巷頂板垂直離層和水平錯(cuò)動(dòng)變形做出雙向的積極響應(yīng),尤其對大斷面、采動(dòng)影響劇烈、頂板軟弱、高應(yīng)力及懸頂面積大等復(fù)雜環(huán)境下的巷道圍巖控制效果突出。(9)依據(jù)桁架錨索在工作狀態(tài)下與頂板巖層緊密貼合的特點(diǎn),建立頂板-桁架錨索組合梁結(jié)構(gòu)力學(xué)模型,分別分析桁架錨索在垂直方向和水平方向?qū)敯迕簬r層垂直撓度和水平錯(cuò)動(dòng)的控制機(jī)理;建立桁架錨索結(jié)構(gòu)承載模型,分析求解出桁架錨索材料的最低抗拉強(qiáng)度和安裝時(shí)的預(yù)緊力,為支護(hù)方案設(shè)計(jì)奠定理論基礎(chǔ)。(10)基于對煤巷采動(dòng)影響劇烈程度分區(qū)的研究,提出了劇烈采動(dòng)煤巷分區(qū)強(qiáng)化控制概念;分析煤巷各區(qū)段圍巖的破壞特征,指出煤巷各區(qū)段相應(yīng)的控制機(jī)理,運(yùn)用數(shù)值建模分析方法與現(xiàn)場實(shí)踐經(jīng)驗(yàn)相結(jié)合,綜合制定各區(qū)段支護(hù)方案;煤巷回采動(dòng)壓影響區(qū):原有支護(hù)整修形成支護(hù)系統(tǒng)→采用桁架錨索控制頂板離層錯(cuò)動(dòng)→補(bǔ)打錨索阻止塑性破壞擴(kuò)展→單體柱提高劇烈采動(dòng)安全儲(chǔ)備;煤巷掘采聯(lián)合動(dòng)壓作用區(qū):提出了以雙向控制為核心的“多支護(hù)結(jié)構(gòu)”重建錨固體的修復(fù)系統(tǒng);煤巷采后靜壓影響區(qū):桁架錨索、單體錨索和錨桿相結(jié)合,多重支護(hù)結(jié)構(gòu)相輔,預(yù)防本工作面回采時(shí)動(dòng)壓影響。(11)對各區(qū)段支護(hù)方案進(jìn)行現(xiàn)場工業(yè)性試驗(yàn),并對采動(dòng)煤巷圍巖控制效果進(jìn)行頂板離層、表面位移監(jiān)測,觀測結(jié)果表明煤巷修復(fù)后,在相鄰工作面劇烈采動(dòng)影響下頂?shù)装遄畲笠平考s138mm,兩幫最大移近量約為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.

【學(xué)位授予單位】：中國礦業(yè)大學(xué)(北京)
【學(xué)位級別】：博士
【學(xué)位授予年份】：2017
【分類號】：TD322;TD353

【參考文獻(xiàn)】

相關(guān)期刊論文 前10條

1 許興亮;田素川;李俊生;茅獻(xiàn)彪;;小紀(jì)汗煤礦工作面頂板破斷結(jié)構(gòu)對巷道礦壓影響規(guī)律研究[J];煤炭學(xué)報(bào);2017年02期

2 年軍;;大采高工作面頂板破斷與失穩(wěn)形式研究[J];煤礦安全;2017年01期

3 王根盛;張偉;朱友恒;王東林;張明鵬;;大斷面煤巷軟厚頂板失穩(wěn)機(jī)理與支護(hù)控制技術(shù)[J];煤礦安全;2016年07期

4 楊登峰;張凌凡;柴茂;李博;白翼飛;;基于斷裂力學(xué)的特厚煤層綜放開采頂板破斷規(guī)律研究[J];巖土力學(xué);2016年07期

5 匡鐵軍;;特厚煤層大采高綜放工作面端部覆巖活動(dòng)規(guī)律研究[J];煤炭科學(xué)技術(shù);2016年06期

6 何富連;張廣超;;大斷面采動(dòng)劇烈影響煤巷變形破壞機(jī)制與控制技術(shù)[J];采礦與安全工程學(xué)報(bào);2016年03期

7 左建平;孫運(yùn)江;王金濤;陳巖;姜廣輝;;大斷面破碎巷道全空間桁架錨索協(xié)同支護(hù)研究[J];煤炭科學(xué)技術(shù);2016年03期

8 于斌;朱衛(wèi)兵;高瑞;劉錦榮;;特厚煤層綜放開采大空間采場覆巖結(jié)構(gòu)及作用機(jī)制[J];煤炭學(xué)報(bào);2016年03期

9 楊勝利;王兆會(huì);孔德中;程占博;宋高峰;;大采高采場覆巖破斷演化過程及支架阻力的確定[J];采礦與安全工程學(xué)報(bào);2016年02期

10 蘇學(xué)貴;宋選民;李浩春;原鴻鵠;李本奎;杜獻(xiàn)杰;;特厚傾斜復(fù)合頂板巷道破壞特征與穩(wěn)定性控制[J];采礦與安全工程學(xué)報(bào);2016年02期

相關(guān)博士學(xué)位論文 前2條

1 吳銳;綜放巷內(nèi)預(yù)充填無煤柱掘巷圍巖結(jié)構(gòu)演化規(guī)律與控制技術(shù)[D];中國礦業(yè)大學(xué);2014年

2 李樹清;深部煤巷圍巖控制內(nèi)、外承載結(jié)構(gòu)耦合穩(wěn)定原理的研究[D];中南大學(xué);2008年

相關(guān)碩士學(xué)位論文 前2條

1 廉常軍;西川煤礦迎采掘巷護(hù)巷煤柱寬度及圍巖控制技術(shù)研究[D];中國礦業(yè)大學(xué);2014年

2 種德雨;利民煤礦迎回采面沿空掘巷圍巖穩(wěn)定控制技術(shù)研究[D];中國礦業(yè)大學(xué);2014年

，

本文編號：1857654