本文選題：雙向大傾角厚煤層 + 綜放開采�。� 參考：《河北工程大學(xué)》2015年碩士論文

【摘要】：隨著我國煤炭資源開采強(qiáng)度的不斷擴(kuò)大,尤其前些年瘋狂的開采,賦存條件良好的礦藏資源越來越少,部分地區(qū)甚至面臨枯竭,大傾角復(fù)雜地質(zhì)條件下的煤層開采技術(shù)有了很大的發(fā)展空間。本論文綜合應(yīng)用理論分析、關(guān)鍵層理論分析、FLAC3D數(shù)值模擬計算等方法,針對大社礦92716雙向大傾角偽傾斜開采工作面實(shí)際地質(zhì)條件,對工作面覆巖結(jié)構(gòu)、覆巖運(yùn)動規(guī)律、頂板控制技術(shù)等進(jìn)行了分析和探討,研究成果對于雙向大傾角煤層安全開采提供理論和現(xiàn)實(shí)依據(jù)。本次研究得出如下結(jié)論:(1)根據(jù)92716工作面的上覆巖層具體情況,運(yùn)用關(guān)鍵層理論,判別出厚度為12.2m的老頂為關(guān)鍵層,垮落帶高度為10-14m之間,裂隙帶高度為38-49m之間,彎曲下沉帶達(dá)到344m;工作面上部冒落帶和裂隙帶發(fā)育高度較下部大。(2)通過模擬92716工作面開采過程中覆巖破壞場演化過程,開始至結(jié)束的整個開采過程中,覆巖與底板隨著工作面的開采,發(fā)生的塑性破壞均較規(guī)律,大煤破壞區(qū)大部分分布在大煤覆巖層,大煤底板只有局部發(fā)生了破壞,開采380m前,大煤直接頂上方高度為6.2m厚的中粒砂巖一直未受到剪切和拉伸破壞,開采至380m時才與附近剪切塑性破壞區(qū)相連,此后直至開采結(jié)束大煤覆巖破壞區(qū)進(jìn)入較穩(wěn)定擴(kuò)展?fàn)顟B(tài)。由覆巖破壞的演化過程得出,在大煤工作面開采至40m處時,關(guān)鍵層發(fā)生剪切破斷,其是老頂?shù)牡谝淮沃芷趤韷?與前面第一章通過材料力學(xué)的方法得到上覆巖層中砂巖的初次垮落步距32.2m基本吻合;當(dāng)大煤工作面采至40m~80m時,隨著工作面不斷推進(jìn),剪切破斷的區(qū)域不斷增大,可以看到開采至80m時,關(guān)鍵層上方巖層發(fā)生過往的剪切破壞,;覆巖破壞區(qū)不斷擴(kuò)大,大煤開采170m時,覆巖破壞區(qū)延伸到大煤覆巖頂部較厚的第二層粉砂巖關(guān)鍵層,大煤破壞場隨著工作面的推進(jìn)沿著走向不斷擴(kuò)展,直到結(jié)束開采。(3)92716工作面的周期來壓步距比較穩(wěn)定,在開采過程中來壓步距在13.9m-16m之間,來壓步距的平均值為14.2m,初次來壓距開切眼平均距離36.3m,雙向大傾角工作面周期來壓強(qiáng)度不明顯,動載系數(shù)的平均值為1.12,表明在使用放頂煤綜采開采大傾角煤層的過程中,采高的增大不會使支架所承擔(dān)的載荷也增大。通過實(shí)測發(fā)現(xiàn)工作面傾斜方向來壓步距不同,首先來壓在工作面上部,后逐漸向工作面下方轉(zhuǎn)移,最后到達(dá)工作面下部,與理論分析基本一致。(4)經(jīng)理論計算確定采面正常使用支架、端頭過渡液壓支架、下巷防滑支架型號為ZF3200/16/24B、ZFG3000/16/24、ZT12000/16/24液壓支架;選擇MG170/410-WD無鏈牽引采煤機(jī);前后選擇SGZ-630/264型可彎曲刮板輸送機(jī)。支架載荷分布不均勻。在支架載荷觀測的過程中,發(fā)現(xiàn)支架前柱的工作阻力大部分工作面小于后柱工作阻力,經(jīng)過實(shí)測和理論分析,造成支架載荷前后支柱不一樣的主要原因有:該工作面俯斜開采,采空區(qū)后方冒落矸石均向工作面方向施加水平推力,造成支架后柱承載了略大于前柱的壓力,根據(jù)來壓步距的大小和規(guī)律,提高支架的初撐力,以求改變近煤壁處圍巖的應(yīng)力狀態(tài),減小端面頂板下沉量,減輕片幫程度,加強(qiáng)圍巖控制,保證來壓期間工作面安全問題。
[Abstract]:With the continuous expansion of the mining intensity of coal resources in China, especially in the past years of crazy mining, the mineral resources with good conditions are less and less, and some areas are even exhausted. The coal mining technology under the complex geological conditions has a great development space. The theoretical analysis of the comprehensive application of this paper, the theoretical analysis of key layers, FLAC 3D numerical simulation calculation and other methods are used to analyze and discuss the overlying rock structure of the working face, the law of overlying rock movement, the roof control technology and so on in view of the actual geological conditions of the 92716 two-way big dip angle pseudo tilt mining face of the Da she coal mine. The research results provide the theoretical and practical basis for the safe mining of the two-way large inclined coal seam. The following conclusions are as follows: (1) according to the specific situation of overlying strata in the 92716 working face, using the key layer theory, the old roof of 12.2m thickness is identified as the key layer, the height of the caving zone is between 10-14m, the height of the fracture zone is 38-49m, the bend subsidence belt reaches 344M; the upper caving zone and fracture zone in the upper part of the working face are higher than the lower part. (2) through simulation 92716. During the whole mining process of overlying rock failure during the working face mining process, the plastic failure of overlying rock and floor with the mining of the working face is more regular, most of the large coal destruction area is distributed in the large coal cladding strata, and the large coal floor is only partially destroyed. Before mining 380m, the height of the direct top of the large coal is above the height. The 6.2m thick medium grain sandstone has not been damaged by shear and tensile. When mining to 380m, it is connected with the nearby shear plastic damage zone, and then until the end of the mining, the failure area of large coal overlying rock enters a stable expansion state. The first periodic pressure of the top is in accordance with the first chapter of the first chapter through the method of material mechanics to get the primary caving distance of the sandstone in the overlying rock. When the large coal face is picked up to 40m~80m, with the continuous advance of the working face, the area of the shear breaking is increasing continuously. It can be seen that the rock stratum above the key layer occurs when the mining to 80m is over. The failure area of the overlying rock enlarges continuously. When the large coal is mined for 170m, the overburden rock failure area extends to the second layer siltstone key layer on the top of the big coal overburden, and the large coal destruction field expands along with the working face along the direction until the end of the mining. (3) the period of the 92716 working surface is more stable and in the process of mining. The average value of the pressure step distance between the step distance and the pressure step distance is 14.2m, the average distance of the opening cutting eye of the initial pressure distance is 36.3m, the cycle pressure strength of the double direction large angle working face is not obvious, the average value of the dynamic load coefficient is 1.12, which indicates that the increase of the height of the mining height will not make the load undertaken by the support in the pass process of the large inclined coal seam with the fully mechanized coal caving mining with the top coal caving mining. It is found that the pressure step is different by measuring the inclination direction of the working face by measuring the working face. First, it is pressed on the top of the working face, and then gradually transferred to the bottom of the working face. Finally, it reaches the lower part of the working face, which is basically consistent with the theoretical analysis. (4) through theoretical calculation, the normal use support of the mining face, the end head transition hydraulic support, and the model of the anti slip support of the lower lane are ZF3200/16 /24B, ZFG3000/16/24, ZT12000/16/24 hydraulic support; select the MG170/410-WD non chain traction shearer; select the SGZ-630/264 type flexural scraper conveyor. The load distribution of the bracket is not uniform. During the load observation of the bracket, it is found that the working resistance of the front column of the bracket is smaller than the rear column. The main reasons for the difference between the supporting pillars before and after the support load are: the inclined mining of the working face and the horizontal thrust of the gangue behind the goaf to the direction of the working face, which causes the rear pillar to carry the pressure slightly larger than the front column, and to improve the initial bracing force according to the size and law of the step distance to change the surrounding rock near the coal wall. The stress state reduces the settlement of the roof face, reduces the extent of the rib, strengthens the surrounding rock control, and ensures the safety of the working face during the pressure.

【學(xué)位授予單位】：河北工程大學(xué)
【學(xué)位級別】：碩士
【學(xué)位授予年份】：2015
【分類號】：TD323;TD355

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