本文選題：綜放開采 + 沿空巷道　； 參考：《中國礦業(yè)大學(xué)(北京)》2017年博士論文

【摘要】：近年來,隨著綜放開采工藝不斷完善及回采設(shè)備大型化、自動(dòng)化程度提高,大型集約化綜放開采已成為我國厚及特厚煤層高產(chǎn)高效安全開采的重要發(fā)展方向。大型綜放開采在實(shí)現(xiàn)厚煤層資源高產(chǎn)高效開采的同時(shí),與之相匹配的綜放回采巷道必然面臨著大斷面、強(qiáng)烈采動(dòng)影響、軟弱厚煤頂?shù)葒鷰r控制難題,加之近年來為響應(yīng)國家建設(shè)資源節(jié)約型礦井的號(hào)召,窄煤柱條件下的綜放大斷面沿空巷道工程越發(fā)普遍。大量現(xiàn)場工程實(shí)踐發(fā)現(xiàn),綜放沿空巷道頂板沿鉛垂方向顯現(xiàn)不對稱下沉破壞,沿水平方向顯現(xiàn)不對稱擠壓變形破壞,且變形破壞的不對稱性在大型綜放開采和區(qū)段窄煤柱條件下趨于惡化,甚至可能引發(fā)惡性冒頂和支護(hù)失效損毀等事故,導(dǎo)致頂板失去控制和煤巷安全性低下。傳統(tǒng)的綜放沿空巷道頂板破壞機(jī)制及相應(yīng)的控制理論與技術(shù)無法有效解決此類巷道圍巖失穩(wěn)問題,有針對性地開展深入系統(tǒng)研究已具有刻不容緩的必要性和緊迫性。本文以王家?guī)X煤礦20103區(qū)段運(yùn)輸平巷為工程背景,基于綜放沿空巷道圍巖性質(zhì)結(jié)構(gòu)和受采動(dòng)影響程度的不對稱差異性,探索頂板煤巖和支護(hù)體不對稱礦壓顯現(xiàn)及其與各影響因素之間的關(guān)聯(lián)性;深究頂板煤巖體的失穩(wěn)準(zhǔn)則和判據(jù)及偏應(yīng)力場時(shí)空演化規(guī)律,闡明頂板不對稱破壞機(jī)制和控制方向;設(shè)計(jì)有效控制頂板不對稱變形破壞的新型錨索桁架結(jié)構(gòu),揭示其礦壓控制作用原理;探究頂板與錨索桁架之間的不對稱調(diào)控關(guān)系和指標(biāo)體系,形成綜放沿空巷道頂板的新型錨索桁架控制系統(tǒng)。取得如下結(jié)論:(1)對綜放沿空巷道頂板變形破壞特征進(jìn)行現(xiàn)場調(diào)研,調(diào)研結(jié)果表明:綜放沿空巷道頂板以巷道中心線為軸沿鉛錘方向和水平方向呈不對稱礦壓顯現(xiàn)特征,沿鉛垂方向,靠煤柱側(cè)頂板嚴(yán)重下沉乃至局部冒漏頂,直接頂與煤柱之間存在滑移、錯(cuò)位、嵌入、臺(tái)階下沉等現(xiàn)象;沿水平方向,頂板巖層水平運(yùn)動(dòng)劇烈,存在圍巖錯(cuò)動(dòng)形成的明顯擠壓破碎帶,并導(dǎo)致了W鋼帶和鋼筋托梁彎曲失效、金屬網(wǎng)撕裂等現(xiàn)象;圍巖變形破壞主要發(fā)生于巷道的靠煤柱側(cè)頂角(45.2%)、靠煤柱側(cè)巷道頂板(25.8%)、煤柱幫中上部(22.6%)3個(gè)位置。(2)采用鉆孔窺視方法觀測頂板內(nèi)裂隙發(fā)育情況,結(jié)果表明:靠實(shí)體煤側(cè)頂板裂隙多以淺部發(fā)育的橫向裂隙及離層和錯(cuò)位為主,而靠煤柱側(cè)頂板裂隙可分為淺部(0~3.0 m范圍)橫向裂隙/離層發(fā)育區(qū)和深部(4.9~12.3 m范圍)走向裂隙發(fā)育區(qū)(局部區(qū)域裂隙完全貫通形成斷裂破碎帶);根據(jù)鉆孔長度、傾斜角度及破碎區(qū)范圍,確定基本頂斷裂位置距采空區(qū)約為5.496~6.847 m。(3)對20103區(qū)段運(yùn)輸平巷圍巖進(jìn)行室內(nèi)力學(xué)實(shí)驗(yàn),評(píng)價(jià)其物理力學(xué)性能。2#煤層單軸抗壓強(qiáng)度為13.89MPa,屬于軟弱煤層,表現(xiàn)出一定的塑性特征;直接頂砂質(zhì)泥巖具有較高的抗壓、抗剪、抗拉強(qiáng)度,巖層相對穩(wěn)定;基本頂粉砂巖單軸抗壓強(qiáng)度達(dá)到142.34MPa,穩(wěn)定性較高;底板泥巖單軸抗壓強(qiáng)度為44.64MPa,遇水泥化。(4)通過理論分析方法研究了綜放沿空巷道基本頂結(jié)構(gòu)特征及其與頂板不對稱破壞的關(guān)系。理論計(jì)算確定了側(cè)向基本頂?shù)膸缀纬叽缗c破斷位置,建立了綜放開采側(cè)向關(guān)鍵塊破斷結(jié)構(gòu)模型,解算得出采動(dòng)影響下沿空巷道頂板巖梁彎矩和撓度表達(dá)式;綜放沿空巷道頂板彎矩和撓度均沿沿巷道中心線呈顯著的不對稱特征,最大彎矩和撓度出現(xiàn)在距煤柱幫1.5m處,該區(qū)域是頂板破壞的關(guān)鍵部位。(5)采用UDEC數(shù)值模擬軟件分析沿空巷道基本頂運(yùn)動(dòng)與巷道圍巖穩(wěn)定性的關(guān)系,探討基本頂破斷位置與沿空巷道不對稱礦壓顯現(xiàn)的定量關(guān)聯(lián)性。淺部0~2.0m范圍內(nèi)靠煤柱側(cè)頂板下沉量明顯大于靠實(shí)體煤側(cè)頂板下沉量,深部2.0~5.0m范圍內(nèi)頂板下沉量自實(shí)體煤側(cè)到煤柱側(cè)呈線性增大趨勢;淺部0~1.5m范圍內(nèi)巖層從兩側(cè)向巷道內(nèi)發(fā)生擠壓運(yùn)動(dòng),0水平位移點(diǎn)由巷道中心處向?qū)嶓w煤側(cè)轉(zhuǎn)移0.9m,深部1.5~6.0m范圍內(nèi)巖層由煤柱側(cè)向?qū)嶓w煤側(cè)發(fā)生運(yùn)動(dòng)。隨著基本頂破斷線與沿空巷道中心線距離減小或者基本頂下沉量增大,煤柱承受的垂直載荷逐漸增大,使得煤柱幫自身承載性能及其對頂板支撐作用明顯弱于實(shí)體煤幫,從而加劇了沿空巷道圍巖結(jié)構(gòu)的不對稱性和頂板變形破壞不對稱性。(6)采用FLAC數(shù)值模擬軟件分析綜放開采不同階段沿空巷道頂板巖層畸變能演化過程。綜放沿空巷道頂板0~3.5m高度范圍內(nèi)偏應(yīng)力第二不變量呈“雙峰狀”分布形態(tài),3.5~11.5m高度范圍內(nèi)偏應(yīng)力不變量呈“單峰狀”分布形態(tài);受本工作面回采期間上覆巖層二次破斷影響,煤柱承載能力降低致使其上方頂板巖層畸變能存儲(chǔ)能力降低,引起偏應(yīng)力第二不變量峰值向?qū)嶓w煤上方頂板轉(zhuǎn)移。隨著煤柱寬度或強(qiáng)度減小,煤柱幫承載能力降低并發(fā)生壓縮變形,造成靠煤柱側(cè)頂板承載能力降低,頂板偏應(yīng)力開始向?qū)嶓w煤上方頂板轉(zhuǎn)移,并誘發(fā)了靠煤柱側(cè)頂板沿垂直方向和水平方向的不對稱位移。(7)基本頂結(jié)構(gòu)回轉(zhuǎn)下沉、松軟煤柱幫、巷道大斷面、支護(hù)不合理等是造成沿空巷道圍巖性質(zhì)結(jié)構(gòu)和應(yīng)力分布不對稱性的主要因素。綜放沿空巷道頂板不對稱破壞災(zāi)變過程如下:相鄰工作面推進(jìn)→基本頂發(fā)生破斷回轉(zhuǎn)下沉運(yùn)動(dòng)→巷道附近區(qū)域煤巖體損傷→巷道開掘促使圍巖性質(zhì)結(jié)構(gòu)和頂板應(yīng)力的不對稱分布→靠煤柱側(cè)煤巖體(頂板、頂角、煤柱幫上部等)局部位移變形→靠煤柱側(cè)頂板煤巖體大范圍破碎或巖層錯(cuò)位、嵌入、臺(tái)階現(xiàn)象→支護(hù)結(jié)構(gòu)載荷增大與不均勻受力→實(shí)體煤側(cè)煤巖體位移變形→大規(guī)模圍巖變形和支護(hù)體破壞→本工作面回采再次激活覆巖結(jié)構(gòu),不對稱變形破壞進(jìn)一步加劇。(8)基于綜放沿空巷道頂板不對稱破壞機(jī)制,提出了以“不對稱式錨梁結(jié)構(gòu)”為核心的綜放沿空巷道調(diào)控系統(tǒng),其主要包括高強(qiáng)錨桿、預(yù)應(yīng)力桁架錨索和不對稱式錨梁結(jié)構(gòu),該調(diào)控系統(tǒng)不但具有控制大范圍塑性破壞、抗剪性能強(qiáng)的優(yōu)點(diǎn),且能對巷道頂板煤巖體變形的不對稱性做出積極響應(yīng)并能對其進(jìn)行有效的控制;研發(fā)的以高強(qiáng)度鋼筋托梁和16#槽鋼托梁為連接構(gòu)件的不對稱錨梁結(jié)構(gòu),具有承壓降載、減垮抗拉、不對稱控制和適應(yīng)頂板水平運(yùn)動(dòng)的特點(diǎn)。(9)結(jié)合20103區(qū)段運(yùn)輸平巷地質(zhì)生產(chǎn)條件進(jìn)行支護(hù)參數(shù)設(shè)計(jì),并提出了現(xiàn)場控制思路。具體支護(hù)措施包括:(1)對頂板進(jìn)行高強(qiáng)錨桿支護(hù),控制圍巖松動(dòng)變形,保證頂板整體性和巷道作業(yè)環(huán)境安全;(2)煤柱幫高強(qiáng)錨桿支護(hù),減少煤柱幫壓縮變形,提高煤柱幫承載能力,降低頂板變形不對稱程度;(3)頂板不對稱式錨梁支護(hù),提高靠煤柱側(cè)頂板承載能力,抑制頂板不對稱下沉和水平擠壓變形;(4)頂角錨索補(bǔ)強(qiáng)加固,提高頂角煤巖穩(wěn)定性,避免局部冒落失穩(wěn);(5)對超前采動(dòng)影響范圍內(nèi)頂板和煤柱進(jìn)行超前加固,進(jìn)一步提高巷道穩(wěn)定性。20103區(qū)段運(yùn)輸平巷現(xiàn)場工程實(shí)踐表明,采用不對稱錨梁支護(hù)系統(tǒng)后,巷道維護(hù)狀況良好,未發(fā)生頂板冒漏現(xiàn)象,頂板不對稱變形破壞得到有效控制,巷道斷面滿足工作面通風(fēng)、運(yùn)輸、行人等要求。
[Abstract]:In recent years, with the continuous improvement of fully mechanized caving mining technology and the large-scale mining equipment, the degree of automation has been improved. The large intensive caving mining has become an important development direction for high yield and high efficiency and safety mining in thick and thick coal seam in China. The roadway must be faced with large section, strong mining influence, weak thick coal top and other surrounding rock control problems. In addition, in response to the call of the country to build a resource saving mine in recent years, the fully mechanized section along the Empty Roadway Project under the condition of narrow coal pillar is more common. A large number of field engineering practice found that the roof of the roadway along the roadway along the space along the vertical direction appears in the vertical direction. Asymmetric subsidence and destruction along the horizontal direction show asymmetric deformation and failure, and the asymmetry of deformation and failure tends to deteriorate under the condition of large caving mining and narrow section of coal pillar, and may even cause accidents such as malignant roof and support failure, which leads to the loss of control of the roof and the low safety of coal roadway. The failure mechanism of the roof and the corresponding control theory and technology can not effectively solve the problem of instability of the surrounding rock of this kind of roadway. It is necessary and urgent to carry out the thorough and systematic research on the roadway. This paper is based on the construction of the roadway of 20103 section of Wangjialing coal mine, based on the structure and mining of the surrounding rock of the roadway along the fully mechanized coal caving. The asymmetrical difference between the dynamic influence degree and the unsymmetrical ore pressure of the roof coal and the support body and the correlation with the influencing factors are explored, the instability criterion and the criterion and the spatio-temporal evolution of the partial stress field of the roof coal rock are studied, and the asymmetric failure mechanism and control direction of the roof are clarified, and the design of the roof asymmetry deformation is designed effectively. The new type of broken anchor cable truss structure reveals the principle of ore pressure control, explores the asymmetric regulation relationship and index system between the roof and the anchor cable truss, and forms a new type of anchor cable truss control system for the roof of the roadway along the goaf. The following conclusions are obtained: (1) field investigation and investigation on the deformation and failure characteristics of the roof of fully mechanized caving roadways The results show that the roof of the roadway along the roadway is asymmetrical ore pressure along the lead and horizontal direction of the roadway along the center line of the roadway. Along the vertical direction of the roadway, the roof of the coal pillar is heavily subsided by the side roof of the pillar, and there is a slip, the dislocation, the step subsidence and so on, and the level of the roof strata along the horizontal direction, and the level of the roof strata along the horizontal direction. Vigorous movement, there is an obvious extrusion crushing zone formed by the dislocation of the surrounding rock, which leads to the failure of the W steel belt and the reinforcement bracket, and the tear of the metal net. The deformation and damage of the surrounding rock mainly occurs in the side corner of the pillar of the coal pillar (45.2%), the roof of the coal pillar side roadway (25.8%), the upper part of the pillar (22.6%) 3 positions. (2) the drilling peep method is adopted. The fracture development in the roof is observed. The results show that the fracture of the roof of the solid coal side is mainly the transverse fissure and the displacement and dislocation in the shallow part, while the fracture of the side roof of the pillar can be divided into the shallow part (0~3.0 m range) and the depth (4.9~12.3 m range) to the fractured zone. According to the length of drilling, the angle of inclination and the scope of broken zone, it is determined that the basic roof fracture position is about 5.496~6.847 M. (3) to carry out the indoor mechanical experiment on the surrounding rock of the 20103 section transportation, and the physical and mechanical properties of the coal seam are evaluated as the single axis anti pressure degree of 13.89MPa, which belongs to the soft coal seam, and shows certain plasticity. The direct top sand mudstone has high compression, shear resistance and tensile strength, and the rock layer is relatively stable; the uniaxial compressive strength of the basic top siltstone is 142.34MPa and the stability is high; the uniaxial compressive strength of the floor mudstone is 44.64MPa and cemented. (4) through theoretical analysis, the basic roof structure characteristics and the top of the fully mechanized caving roadway are studied. The geometric size and breaking position of the lateral basic top are determined by theoretical calculation, and the broken structure model of the key block in fully mechanized mining side is established, and the expression of the bending moment and deflection of the roof of the roadway along the goaf is calculated, and the bending moment and deflection of the top plate along the roadway along the roadway are all along the center line of the roadway. The asymmetry characteristic, the maximum bending moment and deflection appear at the distance from the coal column 1.5m, and this area is the key part of the roof failure. (5) the relationship between the basic roof movement of the roadway along the empty laneway and the stability of the roadway surrounding rock is analyzed by UDEC numerical simulation software, and the quantitative correlation between the basic roof breaking position and the unsymmetrical ore pressure in the roadway along the empty roadway is discussed. The shallow 0~2. The subsidence of the side roof of the coal pillar in the range of 0m is obviously larger than that of the solid coal side roof. The subsidence of the roof in the deep 2.0~5.0m range is linearly increasing from the solid coal side to the pillar side, and the rock stratum in the shallow 0~1.5m ranges from both sides to the roadway, and the 0 horizontal displacement point is transferred from the center of the roadway to the solid coal side 0.9m, and the depth is deep. The rock stratum in the range of 1.5~6.0m is moved by the coal pillar to the solid coal side. The vertical load of the pillar gradually increases with the decrease of the basic top breaking line and the distance of the central line of the roadway along the goaf or the increase of the basic roof subsidence, which makes the bearing performance of the pillar and its supporting effect on the roof weaker than that of the solid coal. The asymmetry of the surrounding rock structure and the roof deformation are asymmetrical. (6) FLAC numerical simulation software is used to analyze the evolution process of the roof rock deformation of the roof of the roadway along the empty laneway in different stages of fully mechanized caving mining. The second invariants of the partial stress in the height range of the roof of the fully mechanized caving roadway are "Shuangfeng", and the height of the 3.5~11.5m is high. The internal partial stress invariants have a "single peak" distribution pattern, and the decrease of the bearing capacity of the coal pillar leads to the decrease of the storage capacity of the upper roof rock distortion and the shift of the peak stress of second unvariable to the top plate above the solid coal. As the width or strength of coal pillar decreases, the coal pillar is reduced by the two breakage of the overlying strata during the recovery period of the working face. The bearing capacity of the supporting capacity is reduced and the compression deformation occurs, which leads to the reduction of the bearing capacity of the side roof of the pillar, the partial stress of the roof begins to transfer to the top roof of the solid coal, and induces the asymmetrical displacement of the side roof along the vertical direction and the horizontal direction. (7) the basic roof structure is slewing down, the soft coal column, the roadway large section and the support are unreasonable, etc. It is the main factor that causes the asymmetry of the nature structure and the stress distribution of the surrounding rock along the goaf. The catastrophe process of the asymmetry failure of the roof of the roadway along the roadway is as follows: the adjacent working face's propulsion, the basic roof, the broken and turning subsidence movement, the damage of the coal and rock in the vicinity of the roadway, and the opening of the roadway impel the nature structure of the surrounding rock and the roof stress. The local displacement and deformation of the coal rock mass (roof, top angle, upper part of coal pillar) of the coal pillar (roof, top angle, upper part of coal pillar) is broken in large scope or the dislocation of the rock layer on the side roof of the coal pillar, and the step phenomenon, the load of the supporting structure and the deformation of the coal and rock mass of the solid coal to the solid coal side, the deformation of the mass surrounding rock and the failure of the support of the supporting body. At the working face, the overlying rock structure is reactivated again, and the asymmetric deformation is further aggravated. (8) based on the asymmetric failure mechanism of the roof of the roadway along the fully mechanized caving, the control system with the core of the "asymmetric anchor beam structure" is put forward, which mainly includes the high strength bolt, the prestressed truss anchor cable and the asymmetric anchor beam structure. The control system not only has the advantages of controlling large plastic damage and strong shearing resistance, but also can respond positively to the asymmetry of roadway roof coal rock deformation and can control it effectively. The unsymmetrical anchor beam structure with high strength steel bar bracket and 16# channel beam as the connecting member has pressure reduction load and collapse resistance. Pull, asymmetric control and adapt to the characteristics of the horizontal movement of the roof. (9) combined with the 20103 section of the roadway geological production conditions for support parameters design, and put forward the idea of field control. Specific support measures include: (1) high strength bolt support to the roof, control of the surrounding rock deformation, to ensure the integrity of the roof and the safety of roadway operation environment; 2) coal pillar to help high strength bolt support, reduce coal pillar compression deformation, improve the bearing capacity of coal pillar and reduce the asymmetry of roof deformation; (3) roof asymmetric anchor beam support, improve the bearing capacity of coal pillar side roof, restrain the roof asymmetry subsidence and horizontal extrusion change; (4) top angle anchorage reinforcement to improve the stability of top angle coal rock, avoid the stability of roof angle coal rock. There is no local caving and instability; (5) advance reinforcement of roof and pillar in the influence range of advanced mining, and further improving the field engineering practice of roadway stability in.20103 section, it shows that the roadway maintenance condition is good, no roof leakage phenomenon occurs, and the asymmetry deformation and failure of roof can be effectively damaged after the use of the unsymmetrical anchor beam supporting system. Control, roadway section to meet ventilation, transportation, pedestrians and other requirements.
【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)】：TD322;TD353

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相關(guān)期刊論文 前10條

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