發(fā)布時(shí)間：2018-05-25 22:05

  本文選題：大采高工作面 + 煤壁破壞機(jī)理　； 參考：《中國礦業(yè)大學(xué)(北京)》2017年博士論文

【摘要】：我國自1978年引進(jìn)國外大采高綜采設(shè)備以來,系統(tǒng)地研究了大采高工作面綜合機(jī)械化采煤方法及其成套裝備技術(shù),并達(dá)到了世界領(lǐng)先水平。然而,煤壁片幫、端面冒頂?shù)裙ぷ髅鎳鷰r穩(wěn)定性問題一直是大采高工作面開采實(shí)踐中最棘手的技術(shù)難題。近年來,隨著裝備制造技術(shù)和生產(chǎn)管理水平的提高,大采高工作面一次采出厚度增大至7-8 m,工作面長(zhǎng)度和開采深度也進(jìn)一步加大,采場(chǎng)采出空間和工作面開采強(qiáng)度顯著增大,因此大面積、大深度的煤壁片幫問題更加突出。大采高工作面煤壁片幫問題嚴(yán)重威脅工作面人員的安全、影響工作面設(shè)備的正常使用和維護(hù)、降低工作面產(chǎn)量和煤礦經(jīng)濟(jì)效益。因此,研究大采高工作面煤壁破壞機(jī)理及其影響因素、尋求工作面煤壁的穩(wěn)定性控制技術(shù),是大采高工作面開采實(shí)踐中亟待解決的關(guān)鍵問題。論文以煤壁破壞機(jī)理及其影響因素、采場(chǎng)系統(tǒng)剛度對(duì)煤壁穩(wěn)定性的影響機(jī)制為研究?jī)?nèi)容,綜合運(yùn)用了理論分析、數(shù)值模擬、實(shí)驗(yàn)室力學(xué)基礎(chǔ)試驗(yàn)、相似模擬試驗(yàn)、工程實(shí)踐等方法,研究了大采高工作面日益突出的煤壁片幫問題。建立了煤壁穩(wěn)定性力學(xué)模型,分析了大采高工作面煤壁破壞機(jī)理,得到了煤壁穩(wěn)定性與頂板載荷、煤壁等效集中力、等效彎矩、護(hù)幫板作用力、護(hù)幫板長(zhǎng)度、煤體內(nèi)聚力、內(nèi)摩擦角等影響因素的關(guān)系;結(jié)合試驗(yàn)手段和數(shù)值計(jì)算,分析了煤中裂隙對(duì)煤體力學(xué)參數(shù)和承載能力的弱化作用,及煤中裂隙對(duì)煤壁破壞特征的影響;根據(jù)工作面推進(jìn)方向上的“采空區(qū)-液壓支架-工作面煤壁”采場(chǎng)系統(tǒng)剛度關(guān)系,建立了采空區(qū)剛度動(dòng)態(tài)演化的數(shù)值模型和支架-煤壁系統(tǒng)剛度力學(xué)模型,進(jìn)行了煤壁穩(wěn)定性三維相似模擬試驗(yàn),研究了采空區(qū)承載性能、煤體GSI、采高、煤壁集中力、煤壁彎矩、支架剛度等對(duì)煤壁穩(wěn)定性的影響機(jī)制;構(gòu)建了基本頂關(guān)鍵巖塊沖擊模型,結(jié)合現(xiàn)場(chǎng)工程實(shí)踐,總結(jié)和提出了煤壁穩(wěn)定性控制原則及煤壁片幫防治措施,取得了以下主要結(jié)論:(1)建立了煤壁穩(wěn)定性力學(xué)模型,采用能量法中基于位移變分原理的Ritz法對(duì)煤壁破壞機(jī)理進(jìn)行了研究,求解了工作面煤體的應(yīng)變能和外力勢(shì)能,得到了工作面前方煤體的位移場(chǎng)和應(yīng)力場(chǎng)分布云圖,其中最大位移和最大應(yīng)力均出現(xiàn)在工作面煤壁上部,即最容易發(fā)生煤壁片幫的位置。(2)根據(jù)莫爾-庫侖屈服條件定義了工作面前方煤體的穩(wěn)定性系數(shù)k:當(dāng)k0時(shí)煤體處于破壞狀態(tài),當(dāng)k=0時(shí)煤體處于極限平衡狀態(tài),當(dāng)k0時(shí)煤體處于穩(wěn)定狀態(tài)。根據(jù)煤體穩(wěn)定性系數(shù)云圖中k0的包絡(luò)范圍,模擬出現(xiàn)場(chǎng)實(shí)踐中3種常見的煤壁片幫形式:煤壁上部片幫、煤壁上下部同時(shí)片幫、煤壁整體片幫;結(jié)合煤壁穩(wěn)定性力學(xué)模型和煤體穩(wěn)定性系數(shù),對(duì)頂板載荷、煤壁等效集中力、煤壁等效彎矩、護(hù)幫板作用力、護(hù)幫板長(zhǎng)度、煤體內(nèi)摩擦角、煤體內(nèi)聚力等煤壁破壞影響因素進(jìn)行了敏感度分析,其中煤體內(nèi)聚力、頂板載荷、煤壁等效集中力、煤壁等效彎矩對(duì)煤壁破壞的影響較大。(3)進(jìn)行了預(yù)制裂隙型煤試件的實(shí)驗(yàn)室力學(xué)參數(shù)試驗(yàn),試驗(yàn)結(jié)果表明:相比于無裂隙的完整試件,含裂隙試件在壓力作用下試件表面產(chǎn)生大量的次生裂隙,試件破碎程度高;節(jié)理跡長(zhǎng)、節(jié)理層間距、節(jié)理連通率對(duì)煤體的力學(xué)參數(shù)弱化效應(yīng)顯著,試件峰值強(qiáng)度當(dāng)量和彈性模量當(dāng)量介于0.5-0.85之間;結(jié)合煤壁穩(wěn)定性力學(xué)模型,煤體損傷變量對(duì)煤壁水平位移及煤壁破壞面積具有較高的敏感度。(4)采用3DEC建立了含節(jié)理煤層大采高工作面煤壁穩(wěn)定性數(shù)值模型,研究了節(jié)理-煤壁方位角α、節(jié)理層間距s、橫向?qū)永韺?duì)煤壁穩(wěn)定性的影響,模擬結(jié)果表明:隨著節(jié)理-煤壁方位角α的增大,煤壁最大位移逐漸減小,但當(dāng)α增大到一定程度時(shí),對(duì)煤壁位移的影響不再明顯,因此在條件允許的情況下,布置工作面時(shí)應(yīng)盡量使工作面傾斜方向與節(jié)理走向相互垂直;隨著節(jié)理層間距s的增大,煤壁位移減小,但當(dāng)s增大到一定程度時(shí),對(duì)煤壁位移的影響不再明顯;當(dāng)煤層中含有橫向?qū)永頃r(shí),煤壁位移僅略微增大,故橫向?qū)永淼挠绊戄^為有限。(5)工作面推進(jìn)方向上的“采空區(qū)-液壓支架-工作面煤壁”采場(chǎng)系統(tǒng)剛度對(duì)采場(chǎng)支承壓力分布規(guī)律、工作面圍巖穩(wěn)定性及破壞具有顯著影響。采用PHASE 2D和FLAC 3D構(gòu)建了采空區(qū)剛度動(dòng)態(tài)演化的數(shù)值模型,完整地模擬出了工作面前后方增壓區(qū)、減壓區(qū)、穩(wěn)壓區(qū)的支承壓力分布特征。數(shù)值模型中,隨著工作面的推進(jìn)和頂板跨距的增大,工作面超前支承壓力、煤體塑性區(qū)寬度、煤壁位移場(chǎng)等表現(xiàn)出先增大后穩(wěn)定的演化規(guī)律。(1)PHASE 2D數(shù)值模擬結(jié)果顯示:當(dāng)采空區(qū)剛度較大時(shí),工作面超前支承壓力較小,支承壓力峰值距煤壁距離縮短,煤壁塑性區(qū)寬度及煤壁最大位移減小,因此提高采空區(qū)剛度有利于增大煤壁穩(wěn)定性;增大煤體GSI、降低工作面采高,工作面超前支承壓力增大,但支承壓力影響范圍減小,塑性區(qū)寬度和煤壁位移降低。(2)FLAC 3D數(shù)值模擬結(jié)果顯示:在工作面傾斜方向上,工作面中部超前支承壓力和煤壁破壞深度較工作面端部更大,因此煤壁破壞具有工作面中部集中效應(yīng);由于采空區(qū)剛度及其對(duì)上覆巖層的承載能力小于實(shí)體煤,故當(dāng)前工作面在靠近已采工作面的端部(較當(dāng)前工作面靠近接續(xù)工作面的端部)的超前支承壓力和煤壁破壞深度均有所增大,因此煤壁破壞具有采空區(qū)影響效應(yīng)。在生產(chǎn)實(shí)踐中,當(dāng)前工作面中部和當(dāng)前工作面靠近已采工作面的端部是煤壁破壞的重點(diǎn)防治區(qū)域。(6)基于彈性地基梁理論建立了支架-煤壁系統(tǒng)剛度力學(xué)模型,求解了工作面前方煤體及液壓支架頂梁的撓曲線方程,計(jì)算結(jié)果表明:工作面前方煤體垂直位移、煤壁垂直位移、支架活柱回縮量與支架剛度呈非線性負(fù)相關(guān),增大支架剛度、提高煤層地基系數(shù),能夠有效降低工作面煤體及支架的變形量;但當(dāng)支架剛度增加到一定程度后,支架對(duì)頂板、煤壁變形的控制作用逐漸減弱。(7)基于煤壁穩(wěn)定性力學(xué)模型和支架-煤壁系統(tǒng)剛度力學(xué)模型,設(shè)計(jì)了大采高工作面煤壁穩(wěn)定性三維相似模擬試驗(yàn),分析了煤壁集中力、煤壁彎矩、支架剛度對(duì)煤壁破壞特征、頂板下沉規(guī)律、煤壁運(yùn)移規(guī)律的作用,揭示了頂板-支架-煤壁系統(tǒng)協(xié)調(diào)變形規(guī)律及支架剛度對(duì)煤壁破壞的影響機(jī)制。試驗(yàn)結(jié)果表明:(1)在煤壁集中力的作用下,當(dāng)頂板千斤頂加載22次時(shí),煤壁發(fā)生整體片幫。片幫時(shí)頂板最大下沉量為27 mm,頂板下沉速率快;煤壁水平位移小,從水平位移開始增長(zhǎng)到煤壁發(fā)生片幫所經(jīng)歷的時(shí)間間隔為200 s,煤壁水平位移變化速率快;煤壁片幫具有突發(fā)性,即煤壁在短時(shí)間內(nèi)累積了較小的垂直位移和水平位移即可觸發(fā)煤壁片幫。(2)在煤壁彎矩的作用下,當(dāng)頂板千斤頂加載26次時(shí),煤壁發(fā)生整體片幫。片幫時(shí)頂板最大下沉量為35 mm,頂板下沉速率較快;煤壁水平位移開始增長(zhǎng)到煤壁片幫發(fā)生所經(jīng)歷的時(shí)間為500 s,煤壁水平位移變化速率較快;煤壁片幫具有一定的突發(fā)性。(3)當(dāng)支架剛度較小時(shí),頂板千斤頂加載81次后,煤壁發(fā)生整體片幫。支架增阻期間頂板下沉速率較慢,但支架卸載期間頂板下沉顯著,并呈現(xiàn)階梯式下沉規(guī)律,下沉梯度為5 mm,片幫時(shí)頂板最大下沉量為47 mm;煤壁水平位移大,從水平位移開始增長(zhǎng)到煤壁發(fā)生片幫的時(shí)間間隔為1200 s,煤壁水平位移變化速率較慢;在支架的支護(hù)作用下,煤壁在較長(zhǎng)的時(shí)間內(nèi)累積了較大的垂直位移和水平位移,最終引發(fā)煤壁片幫。(4)支架剛度增大后,當(dāng)頂板千斤頂加載86次時(shí),煤壁僅發(fā)生局部片幫。支架增阻期間頂板下沉緩慢,并呈現(xiàn)階梯式下沉規(guī)律,下沉梯度為3-3.5 mm,片幫時(shí)頂板最大下沉量為35 mm;煤壁水平位移量較大,水平位移開始增長(zhǎng)到煤壁發(fā)生片幫的時(shí)間間隔為1400 s,煤壁水平位移變化速率緩慢,煤壁片幫現(xiàn)象有所緩解。(5)支架剛度進(jìn)一步增大后,頂板千斤頂加載87次時(shí)未發(fā)生煤壁片幫現(xiàn)象,工作面僅出現(xiàn)局部破碎情況。頂板下沉慢,階梯下沉梯度為2 mm,片幫時(shí)頂板最大下沉量為25.55 mm;煤壁水平位移小,水平位移變化速率慢。因此當(dāng)支架剛度足夠大時(shí),煤壁在有限的時(shí)間內(nèi)所積累的垂直位移和水平位移較小,不足以觸發(fā)煤壁片幫。(8)建立了周期來壓期間基本頂關(guān)鍵巖塊沖擊力學(xué)模型,確定了直接頂、支架、工作面煤體的剛度對(duì)頂板載荷、煤壁集中力、煤壁彎矩的影響作用,總結(jié)和提出了緩解頂板載荷、降低煤壁集中力、控制基本頂破斷巖塊回轉(zhuǎn)的煤壁穩(wěn)定性控制原則;根據(jù)王莊煤礦8101大采高工作面支架工作阻力利用率不高的現(xiàn)狀,指出了支架升柱時(shí)間短、供液不充分、初撐力不足的問題,提出了增大液壓支架剛度和初撐力、提高護(hù)幫板使用率、工作面煤壁注漿、優(yōu)化工作面回采工藝的的煤壁片幫防治措施,取得了良好的煤壁片幫控制效果。
[Abstract]:Since the introduction of foreign large mining high mechanized mining equipment in China in 1978, the comprehensive mechanized coal mining method and complete set of equipment technology for large mining face have been systematically studied, and the world leading level has been reached. However, the stability of the surrounding rock of the working face, such as the coal wall section and the end face, has been the most difficult technique in the mining practice of the large mining face. In recent years, with the improvement of equipment manufacturing technology and production management level, the thickness of the mining face of large mining height increased to 7-8 m, the length of working face and mining depth increased further, the mining intensity of mining area and working face increased significantly, so the large area and large depth of coal wall slaving problem became more prominent. The problem of coal wall section of working face seriously threatens the safety of staff in working face, affects the normal use and maintenance of working face equipment, reduces the output of working face and the economic benefit of coal mine. Therefore, the study on the mechanism of coal wall failure and its influencing factors and the stability control technology of coal wall in the working face are the mining face of large mining face. In this paper, the mechanism of coal wall failure and its influencing factors, the influence mechanism of the stiffness of the stope system on the stability of the coal wall are studied, and the theoretical analysis, numerical simulation, the laboratory mechanics basic test, the similar simulation test and the engineering practice are used to study the increasingly prominent working face of the large mining height. The mechanical model of coal wall stability is established, and the failure mechanism of coal wall is analyzed. The relationship between the stability and roof load of the coal wall, the equivalent concentrated force of coal wall, the equivalent bending moment, the force of the retaining wall, the length of the retaining plate, the cohesion of the coal body, the internal friction angle and so on; The weakening effect of the cracks on the mechanical parameters and bearing capacity of coal body and the influence of the cracks in coal on the damage characteristics of the coal wall are analyzed, and a numerical model of the dynamic evolution of the stiffness of the goaf and the stiffness of the support coal wall system is established according to the stiffness relationship of the mining area system of the "goaf - hydraulic support - working face coal wall" in the direction of the working face. A three dimensional simulation test of the stability of coal wall is carried out by the degree of mechanical model. The bearing performance of the goaf, the coal body GSI, the coal wall concentration force, the coal wall bending moment, the rigidity of the coal wall, and so on, the impact mechanism of the coal wall stability are studied. The impact model of the key rock block is constructed and the stability control of the coal wall is concluded and put forward. The main conclusions are as follows: (1) the mechanical model of coal wall stability is established. The failure mechanism of coal wall is studied by Ritz method based on the principle of displacement and variation in energy method. The strain energy and external force potential of coal face are solved, and the displacement field and stress field of coal body in front of the working face are obtained. The maximum displacement and maximum stress appear on the top of the coal wall on the working face. (2) according to the Mohr Coulomb yield conditions, the stability coefficient of the coal body in front of the working face is defined k: when the coal body is in the state of failure when K0, when the coal body is in the limit equilibrium state, when K0 is in the coal body. According to the envelope range of K0 in the coal body stability coefficient cloud, 3 kinds of common coal wall sheet help forms in field practice are simulated: the upper part of the coal wall, the upper and lower part of the coal wall and the whole wall of the coal wall, combined with the stability mechanics model of coal wall and the number of coal stability system, the load of the roof, the equivalent concentrated force of the coal wall and the equivalent of the coal wall. The sensitivity analysis of the bending moment, the force of the retaining plate, the length of the retaining plate, the friction angle of the coal body and the cohesion of coal in the coal wall, including the cohesion of the coal, the load of the roof, the equivalent concentrated force of the coal wall and the equivalent bending moment of the coal wall have great influence on the failure of the coal wall. (3) the Laboratory mechanical parameters of the prefabricated crack briquettes are tested. The test results show that a large number of secondary cracks are produced on the surface of the specimen under pressure action compared to the complete specimen without fissure, the fracture degree of the specimen is high, the length of the joint, the spacing of the joint layer, the joint rate of joint on the mechanical parameters of the coal body is significant, and the peak strength equivalent and the modulus of elasticity of the specimen are in the 0.5-0.85. In connection with the mechanical model of coal wall stability, the damage variable of coal body has high sensitivity to the horizontal displacement of coal wall and the damaged area of coal wall. (4) a numerical model of the stability of coal wall with a large mining face with joint coal seam is established by 3DEC. The influence of the azimuth angle of the joint - coal wall, the spacing of the joint layer of S, and the lateral bedding on the stability of the coal wall The simulation results show that the maximum displacement of coal wall decreases with the increase of the azimuth of the joint - coal wall azimuth, but when the alpha increases to a certain extent, the effect of the coal wall displacement is no longer obvious. Therefore, when the condition is allowed, the working face should be arranged as far as the inclined direction of the working face is perpendicular to the joint direction; with the spacing of the joint layer s The displacement of coal wall decreases, but when the s increases to a certain extent, the influence of the coal wall displacement is no longer obvious; when the coal seam contains transverse bedding, the displacement of the coal wall only slightly increases, so the lateral bedding is more limited. (5) the stope stiffness of the "goaf - hydraulic support - working face coal wall" in the direction of the working face Bearing pressure distribution law, the stability and failure of surrounding rock has significant influence. A numerical model of dynamic evolution of the stiffness of the goaf is constructed by PHASE 2D and FLAC 3D, which completely simulates the distribution characteristics of the support pressure in the turbocharging area, the pressure reducing area and the stable pressure area behind the work. The increase of distance, the overbearing pressure of the working face, the width of the coal body plastic zone and the displacement field of the coal wall first increase and then the stable evolution law. (1) the results of PHASE 2D numerical simulation show that, when the rigidity of the goaf is large, the overbearing pressure of the working face is smaller, the peak value of the supporting pressure is shorter than the coal wall distance, the width of the plastic zone of the coal wall and the largest coal wall. When the displacement is reduced, the rigidity of the goaf is beneficial to increase the stability of the coal wall; increase the coal body GSI, reduce the height of the working face, increase the front support pressure of the working face, but reduce the influence range of the support pressure, the width of the plastic zone and the coal wall displacement. (2) the FLAC 3D numerical simulation results show that the middle of the working face is ahead of the working face in the direction of the face. The support pressure and coal wall failure depth is greater than the end of the working face, so the coal wall failure has the central effect in the middle of the working face; because the rigidity of the goaf and the bearing capacity of the overlying rock are less than the solid coal, the current working face is near the end of the working face near the present working face near the end of the working face. Pressure and coal wall failure depth have increased, so coal wall failure has the effect of goaf. In the production practice, the current working face and the present working face near the end of the working face are the key prevention areas of the coal wall failure. (6) based on the elastic foundation beam theory, the stiffness mechanics model of the support coal wall system has been established, and the solution is solved. The calculation results show that the vertical displacement of coal body in front of the working face, the vertical displacement of coal wall, the nonlinear negative correlation of the back shrinkage of the support column and the stiffness of the support, increase the stiffness of the support and the coefficient of the foundation of the coal seam, and can effectively reduce the deformation of the coal body and the support of the working face; but when the coal body and the support are reduced, the deformation of the coal face and the support can be effectively reduced. After the stiffness of the support is increased to a certain degree, the control of the support to the roof and the deformation of the coal wall gradually weakened. (7) based on the mechanical model of the stability of the coal wall and the stiffness model of the support - coal wall system, a three-dimensional similarity simulation test of the stability of the coal wall in a large mining face was designed, and the coal wall concentration force, the coal wall bending moment, and the stiffness of the support on the coal wall were analyzed. The bad characteristics, the roof subsidence law and the action of the coal wall movement, reveal the mechanism of the coordination deformation of the roof support coal wall system and the influence mechanism of the support stiffness on the failure of the coal wall. The test results show that (1) the coal wall has a whole wall when the top plate jack is loaded 22 times under the action of the coal wall concentrated force. For 27 mm, the roof subsidence rate is fast, the horizontal displacement of coal wall is small, the time interval from the horizontal displacement to the coal wall is 200 s, the horizontal displacement of coal wall is fast, and the coal wall section is sudden, that is, the small vertical displacement and horizontal displacement of coal wall can trigger the coal wall section. (2) coal Under the action of wall bending, when the roof jack is loaded 26 times, the coal wall takes a whole slice. The maximum subsidence of the roof is 35 mm, the roof subsidence rate is faster. The coal wall horizontal displacement begins to grow to the coal wall section, the time is 500 s, the horizontal displacement of the coal wall is faster, and the coal wall section has a certain burst. (3) When the support stiffness is small, the top plate jack is loaded 81 times after the roof is loaded. The top plate sinking rate is slower, but the roof subsidence rate is slow during the support increase, but the roof subsidence is remarkable, and the downward gradient is 5 mm, the maximum subsidence of the roof is 47 mm, the horizontal displacement of the coal wall increases from the horizontal displacement. The time interval between the coal wall and the coal wall is 1200 s, and the change rate of the horizontal displacement of the coal wall is slow. Under the support of the support, the coal wall accumulates a large vertical displacement and horizontal displacement in a long time. (4) when the stiffness of the support is increased, the coal wall only takes part in the local section when the jack is loaded 86 times. The roof subsidence is slow and the subsidence gradient is 3-3.5 mm, the maximum subsidence of the roof is 35 mm, the horizontal displacement of the coal wall is larger, the horizontal displacement begins to increase to 1400 s, the change rate of the coal wall horizontal displacement is slow, and the ganging phenomenon of coal wall is relieved (5). (5 ) when the stiffness of the bracket is further increased, there is no coal wall segment phenomenon when the roof jack is loaded 87 times. Only local breakage occurs in the working face. The roof subsidence is slow, the gradient of the staircase is 2 mm, the maximum subsidence of the roof is 25.55 mm, the horizontal displacement of the coal wall is small and the horizontal displacement is slow. So when the stiffness of the support is large enough, coal The vertical displacement and horizontal displacement accumulated in a limited time are not enough to trigger the coal wall section. (8) the impact mechanical model of the key rock block is established during the period of periodic pressure. The influence of the stiffness of the direct top, the support and the coal face on the roof load, the coal wall concentration force and the coal wall bending moment is determined. To solve the roof load, reduce the coal wall concentration force, control the principle of the stability control of the coal wall of the broken rock block at the top of the top, and according to the present situation of the low utilization rate of the working resistance of the support in the 8101 large mining face of Wang Zhuang coal mine, points out the problem that the supporting column time is short, the supply fluid is insufficient and the initial brace force is insufficient, and the stiffness of the hydraulic support and the initial support are raised. Force, improve the utilization rate of the retaining wall, the coal wall grouting in the working face, optimize the control measures of the coal wall section of the coal wall, and obtain the good control effect of the coal wall section.
【學(xué)位授予單位】：中國礦業(yè)大學(xué)(北京)
【學(xué)位級(jí)別】：博士
【學(xué)位授予年份】：2017
【分類號(hào)】：TD823;TD323

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本文編號(hào)：1934809