Fracture propagation and evolution law of indirect fracturing in the roof of broken soft coal seams

Indirect fracturing in the roof of broken soft coal seams has been demonstrated to be a feasible technology.In this work,the No.5 coal seam in the Hancheng block was taken as the research object.Based on the findings of true triaxial hydraulic fracturing experiments and field pilot under this technology and the cohesive element method,a 3D numerical model of indirect fracturing in the roof of broken soft coal seams was established,the fracture morphology propagation and evolution law under different conditions was investigated,and analysis of main controlling factors of fracture parameters was conducted with the combination weight method,which was based on grey incidence,analytic hierarchy process and entropy weight method.The results show that“士”-shaped fractures,T-shaped fractures,cross fractures,H-shaped fractures,and“干”-shaped fractures dominated by horizontal fractures were formed.Different parameter combinations can form different fracture morphologies.When the coal seam permeability is lower and the minimum horizontal principal stress difference between layers and fracturing fluid injection rate are both larger,it tends to form“士”-shaped fractures.When the coal seam permeability and minimum horizontal principal stress between layers and perforation position are moderate,cross fractures are easily generated.Different fracture parameters have different main controlling factors.Engineering factors of perforation location,fracturing fluid injection rate and viscosity are the dominant factors of hydraulic fracture shape parameters.This study can provide a reference for the design of indirect fracturing in the roof of broken soft coal seams.

Overburden fracture evolution laws and water-controlling technologies in mining very thick coal seam under water-rich roof

Considering the danger of water inrush in mining very thick coal seam under water-rich roof in Majialiang Coal Mine,the universal discrete element(UDEC)software was used to simulate the overburden fracture evolution laws when mining 4#coal seam.Besides,this study researched on the influence of face advancing length,speed and mining height on the height of the water flowing fractured zones(HWFFZ),and analyzed the correlation of face advancing length and change rules of aquifer water levels and goaf water inflow.Based on those mentioned above,this research proposed the following water-controlling technologies:draining the roof water before mining,draining goaf water,reasonable advancing speed and mining thickness.These water-controlling technologies were successfully used in the feld,thus ensured safely mining the very thick coal seam under water-rich roof.

Study on the Fissure Rate in the Roof Strata of Excavated Coal Seams

In the past decades,the intense excavation of coal resources in China has induced a series of severe waterrelated disasters or problems,such as water burst accidents,severe groundwater depletion and contamination.All of these problems were caused by the structural changes of roof rocks or bed rocks induced by coal excavation.The structure of collapsed roof rocks is generally classified into three zones：falling zone.

Control mechanism of a cable truss system for stability of roadways within thick coal seams

Cable truss systems have been widely applied in roadways with complicated conditions, such as the large cross-sections of deep wells, and high tectonic stress. However, they are rarely applied to roadways with extremely thick coal seams because the control mechanism of the system for the deformation of the roof and the separation between coal rock segments is not completely understood. By using the relationship between the support system and the roof strata, a mechanical model was established to calculate the deformation of the roof in a thick coal seam with bedding separation under different support conditions: with an anchor truss support and without support. On this basis, the research was used to deduce a method for computing the minimum pre-tightening forces in the anchor truss, the maximum amounts of subsidence and separation with, and without, anchor truss support under the roof, and the maximum subsidence and the decreasing amounts of the separation before and after adopting the anchor truss. Additionally, mechanical relationships between the minimum pre-tightening force and the anchoring force in the anchor were analyzed. By taking a typical roadway with thick coal roof as an example, the theoretical results mentioned above were applied in the analysis and testing of a roof supporting project in a roadway field to verify the accuracy of the theory: favorable experimental results were achieved. In addition, the relationships among other parameters were analyzed, including the minimum pre-tightening forces applied by the anchor truss, the angle of inclination of the anchor cable, and the array pitch. Meanwhile, the changing characteristics of the amounts of roof separation and subsidence with key parameters of the support system(such as array pitch, pre-tightening force, and inclination angle) were also analyzed. The research results revealed the acting mechanism of the anchor truss in control of roadway stability with a thick coal seam, providing a theoretical basis of its application in coal mining.

Controlling Effects of Surrounding Rock to Coal Seam Gas Occurrence

Surrounding rock of coal seam was one of the important factors to gas occurrence. The coal seam gas occurrence was studied by the index of roof strata thickness or sand content rate;we found that there were certain shortcomings. In order to reasonably evaluate the influence of coal seam surrounding rock on gas occurrence in Panji mining area, we quantitatively evaluated the effect of coal seam surrounding rock on gas occurrence by influence coefficient of roof strata thickness, and built six mathematical models of the variational gas content in the mining area which is divided into six gas geological units. The results shows that the coal seam gas content is mainly influenced by 20 mroof strata in each gas geological unit, the gas content presents the tendency of increase, and with the influence coefficient of strata thickness increases, they exist a significant linear relationship.

极近距离煤层开采无煤柱自成巷控制方法研究

极近距离煤层传统长壁开采,受上层遗留煤柱和采空区垮落等因素的影响,下煤层开采过程中应力环境复杂、矿压显现剧烈,易诱发巷道围岩大变形。为解决上述问题,提出了极近距离条件下巷道定向切顶–约束高强支护无煤柱自成巷控制方法。通过顶板定向预裂切顶,主动改变顶板悬臂结构,切断采空区向巷道顶板的应力传递。充分利用矿山压力和岩体碎胀特性,取消煤柱留设,结合高强支护加强巷道顶板整体性,共同实现切顶自成巷。建立了极近距离煤层开采覆岩结构模型,计算了下煤层切顶自成巷巷旁支护阻力。以典型极近距离煤层为工程背景,开展了不同开采方法的数值试验对比研究,结果表明,提出的自成巷控制方法使巷道围岩应力降低59.8%,巷道顶板变形减少70.8%,并明确了极近距离煤层开采无煤柱自成巷控制机理。在此基础上,开展了典型极近距离煤层工程设计及现场应用研究,结果表明该方法有效降低了矿压显现程度,保证了自成巷的安全稳定控制。

大采高综采工作面切顶留巷矸石帮变形控制技术研究

为解决厚煤层大采高综采工作面切顶留巷矸石帮围岩控制难题,首先利用理论分析建立了不同挡矸支护条件下矸石帮位移变化的力学公式,以此为基础提出矸石帮变形控制双策略,减弱矸石帮变形压力和增强矸石帮的抗变形能力;其次利用现场实验设计两种类别共4组挡矸支护方案,探讨不同支护方案的矸石帮控制效果,验证理论研究成果。结果表明:①增强矸石帮侧顶板支护强度,能减弱矸石帮变形压力;②增架卡兰个数、增加卡兰预紧力、增加卡兰个数、增强U型钢之间的摩擦力等能增加U型钢滑移初始滑移应力,增强矸石帮的抗变形能力;③挡矸支护控制效果为:2副卡兰+单体支柱<2副卡兰+单元支架<2副卡兰+单元支架+呛柱<3副卡兰+单元支架。该研究可为同类条件下切顶留巷矸石帮变形控制提供参考。

深部顶板夹煤层巷道变形破坏分析及其控制

为了探究深部顶板夹煤层软岩巷道大变形破坏问题,以平煤六矿三水平戊二胶带运输巷为工程背景,开展夹煤层巷道围岩破坏分析及控制对策的研究。首先,构建原支护体系组合拱力学模型,推导出直墙半圆拱形巷道稳定判别式;其次,提出“锚网索-梁-注浆-组合砂浆锚索”联合返修控制对策,即沿戊8煤层顶板进行巷道扩刷,采用锚网索、梯子梁完成基本支护,对围岩进行深浅孔注浆加固、顶板及两帮关键部位布置组合砂浆锚索等加强支护,并通过数值模拟分析返修前后应力、位移场分布特征;最后,开展工业性试验,并监测矿压。结果表明:两帮及顶底板最大收敛量分别为93、112 mm,后期变形速率均低于1 mm/d,围岩变形得到有效控制,返修效果良好。

淮南矿区煤层顶板分段压裂水平井抽采技术及效果研究

淮南矿区为典型高瓦斯矿区,煤层碎软、渗透率低、瓦斯含量偏高、抽采难度大,为探讨地面煤层气顶板分段压裂水平井在矿区的技术可行性与瓦斯治理效果,在分析矿区主要煤层13-1煤储层特征基础上,采用应力解除法进行了煤层三向地应力测试,结果显示三向应力场类型主要为σ_(h,max)>σ_(v)>σ_(h,min),具有实施顶板分段压裂水平井技术的充分条件;利用MFrac Suite软件分别模拟了水平段距离煤层1、3、5 m时的压裂缝参数,压裂缝半长最大107.33 m、最小89.47 m,具有理想的压裂效果,说明顶板分段压裂水平井在淮南矿区具有比较好的地质适应性与可行性。以潘一煤矿13-1煤层“L”型顶板分段压裂水平井CBM01井为研究对象,采用井下钻孔检测与数值模拟等手段综合分析了瓦斯治理效果,结果显示CBM01井抽采415 d即显著降低了煤层瓦斯压力与瓦斯含量,距离水平井50、65 m处瓦斯压力由6.4 MPa分别降至2.6、2.7 MPa,降低幅度均超过55%,水平段两侧各15~20 m范围内瓦斯含量由13.5 m^(3)/t降至最大9.11 m^(3)/t、最小6.92 m^(3)/t,平均7.92 m^(3)/t,约10 m范围均降至8 m^(3)/t以下。最后,采用数值模拟方法预测了CBM01井抽采10 a的产气效果及瓦斯治理效果,气井抽采10 a累计产气约272.08×10^(4)m^(3),水平段倾向单侧约150 m范围内气含量均降至8 m^(3)/t以下、压力均降至3 MPa以下。综合研究结果表明,煤层顶板分段压裂水平井技术在淮南矿区瓦斯治理方面均具有较大的优势和应用效果。

托顶煤巷道易片冒顶板变形机理与超前导管预注浆控制技术研究

目的 为探究厚煤层易片冒区域托顶煤巷道顶板变形破坏机理,并对易片冒区托顶煤巷道顶板进行有效控制,方法 以常村煤矿2701工作面皮带运输巷为例,通过等截面梁理论分析托顶煤巷道顶板破断成因,研究托顶煤巷道顶板的稳定性与巷道宽度和顶煤强度的关系。通过极限平衡准则得到巷帮位移的计算公式,证明托顶煤巷道具有帮顶协同变形机制。建立托顶煤巷道数值模型,研究巷道埋深、侧压系数、顶煤厚度和顶煤强度与托顶煤巷道围岩稳定性的关系。结果 结果表明:巷道埋深越大,其对托顶煤巷道两帮稳定性的影响越大;侧压系数的增加对巷道两帮和底鼓量的影响更大,对顶板影响较小;顶板下沉量与顶煤厚度呈正相关,底鼓量则与顶煤厚度呈负相关;顶煤强度的增加使巷道顶板下沉量与两帮移近量均近似呈线性减小趋势,底鼓量基本保持不变。结论 通过幂律型流体柱形渗透注浆扩散理论得到超前导管注浆半径,由此提出托顶煤巷道超前导管预注浆帮顶联合控制技术,成功控制易片冒托顶煤巷道顶板冒落问题,可为类似地质条件的托顶煤巷道顶板控制提供借鉴。

薄煤层灰岩顶板巷道围岩稳定性控制

为合理控制薄煤层灰岩顶板巷道稳定性,避免过度支护引起巷道顶板失稳。以鲁西某矿薄煤层灰岩顶板巷道为研究对象,运用数值模拟、理论计算等研究方法,研究了不同类型灰岩顶板采场应力分布特征、围岩矿压显现规律及其塑性破坏形态特征,灰岩临界安全厚度影响性分析。研究结果发现:实体巷道的应力环境较好,应力集中程度相对较低,而沿空巷受2次采动影响范围更大、应力集中程度更高。煤矿巷道围岩在不同应力环境下需要不同程度的支护,主要集中在加强巷道两帮的支护和在特定情况下加强超前影响范围内的支护。

倾斜煤层厚硬顶板切顶留巷关键参数优化研究

为解决回采巷道倾斜厚硬顶板应力集中问题,降低顶板覆岩突然垮落的冲击扰动风险及提高倾斜厚硬顶板切顶留巷围岩控制效果,以长城六矿1301N运输巷切顶留巷为工程背景,综合采用理论分析、数值模拟及现场监测的研究方法开展倾斜煤层厚硬顶板切顶留巷关键参数优化研究。首先,基于岩石碎胀系数、砌体梁理论及现场地质资料,推导计算了1301N运输巷切顶高度和切顶角度的理论最小值。其次,为进一步验证并优化理论推导结果,利用PFC^(2D)数值软件建立了1301N工作面切顶留巷模型,采用控制变量法分析了不同切顶高度和切顶角度下留巷顶板应力、位移及组构张量响应特征。分析结果表明:切顶高度与卸压效果之间存在底数大于1的对数增长关系,即随着切顶高度增加,卸压效果逐渐增强但其增长幅度逐渐减小;切顶角度与卸压效果之间存在“S”型增长关系,即随着切顶高度增加,卸压效果先增强后减弱。卸压参数与卸压效果之间的变化规律揭示了切顶高度选择存在最适值,切顶角度选择存在最优解。综合理论分析、经济收益及数值模拟结果,确定长城六矿1301N运输巷倾斜厚硬顶板切顶留巷过程中最适切顶高度为13 m,最优切顶角度为10°。现场监测获得该切顶参数下围岩控制效果较好,留巷围岩变形量较小,可有效满足下一工作面开采需求,验证了优化切顶参数的有效性,为类似地质条件切顶留巷参数的选取提供理论依据和实践参考。

大采高工作面顶板结构模型及液压支架支护阻力计算

针对大采高工作面覆岩垮落范围大,易出现液压支架稳定性差、损坏率高以及支架压死的问题,采用物理模拟、理论分析以及工程实践的方法,对大采高工作面顶板破断结构模型与支架合理支护阻力计算方法进行了研究。结果表明:随着大采高工作面推进,受采空区空间大的影响,顶板将呈现“组合悬梁-非铰接顶板-铰接顶板”结构特征;支架与围岩相互作用体系由支架-组合悬梁结构-非铰接顶板结构-铰接顶板结构组成,揭示了各结构间的相互作用关系并确定了大采高工作面支架工作阻力的计算公式,支架支护阻力应适应覆岩结构失稳运动的变化,承载结构自身重量以及运动产生的附加载荷;结合大同矿区晋华宫煤矿大采高工作面开采条件,计算了工作面支架合理支护阻力并进行了支架选型,矿压监测显示,选择的ZZ13000/28/60型支撑掩护式液压支架能满足顶板控制的要求,保障了大采高工作面的安全回采。

近浅埋多层厚硬顶板大采高综放工作面矿压显现规律研究

针对榆神矿区近浅埋多层厚硬顶板条件下,“大周期”来压支架立柱下缩量大、煤柱应力集中系数高、稳定性差等强矿压问题,采用了数值计算、微震和煤体应力监测等技术手段,对覆岩破断运动及工作面来压规律进行了分析研究,结果表明:覆岩破断运动上,多层厚硬顶板较大的“垂向离层空间”和“自下而上”的顺次交替垮落孕育了顶板“高+低层位组合破断”的时空基础;“高低层位顶板组合破断”是“大周期”来压增强的关键诱因;同时,煤柱应力的时空演化受控于多层厚硬顶板的破断垮落。矿压显现上,工作面“大周期”立柱下缩量平均可达945 mm,持续距离12.10 m,并可由2~5次连续强来压组成;巷道煤柱应力集中系数达到1.80~2.30,影响范围为工作面前方126 m至后方471.3 m,具有明显的“高强度、大范围、长上升期”特点。针对该条件下的强矿压防治,分别提出了“弱化低位顶板、优化破断结构”和“弱化高位顶板、改变破断次序”的技术思路,并取得的了较好的应用效果。

复合再生顶板条件下煤体固化与顶板强化技术研究及应用

为提升复合再生顶板工作面的安全性和回采效率,在平煤十二矿己16-17-31020复合再生顶板工作面开展了煤体注水固化、顶板注浆强化和管缝锚杆顶板强化试验,探索了复合再生顶板工作面矿压显现规律,评价了加固措施的支护效果。研究结果表明:加固措施实施后,工作面完整度大幅提升,降低了顶板冒落和压架等事故风险;平均来压步距由22.15 m增加至29.92 m,顶板压力由25~31 MPa降到19~26 MPa。顶板下沉量降低了50.4%,保证了工作面的正常安全回采,同时进尺速率提升了16.7%。煤体固化-顶板强化措施取得了良好的效益,可为同类型复合再生顶板工作面提供借鉴。

下峪口煤矿23306下工作面回采巷道合理错距研究

近距离煤层采空区下回采巷道合理错距的确定对巷道安全高效掘进意义重大。以下峪口煤矿为工程背景,采用理论分析及现场试验的方法,对23306下进顺采空区下回采巷道合理错距进行研究。得出下峪口煤矿23306下进风顺槽合理留设错距为22 m,根据23306下进顺错距留设情况提出支护方案并根据巷道顶底板及两帮位移变化情况进行监测验证,监测发现巷道各测点顶底板及两帮的移近量均小于450 mm,大多时间内两帮的移近量小于350 mm,煤柱帮的深基点和浅基点的位移变化量基本一致,且随着巷道的掘进,未出现较大变形,说明采用22 m错距可实现围岩变形稳定,研究可为类似条件巷道安全高效生产提供经验和参考。

分段多簇密切割压裂技术在淮南矿区煤层气抽采中的应用

针对淮南矿区碎软低渗煤层瓦斯抽采技术难题,以13-1煤层为例,在分析煤层及其顶底板概况的基础上,采用煤层顶板分段压裂水平井技术,充分利用水平井水平段,单段分3簇射孔,簇间距25 m,实现分段多簇密切割体积压裂改造。微地震监测结果表明,单段裂缝长度102.5~211.4 m,缝高17.3~27 m,裂缝主体向下延伸,实现了井筒与煤层的沟通。排水采气效果表明,最高产气量1519 m^(3)/d,取得了良好的产气效果。研究成果为淮南矿区煤层气地面预抽和瓦斯区域治理提供了工程借鉴。

煤层顶板高位大直径定向钻孔瓦斯抽采技术研究

在煤层瓦斯抽采工程中,在钻场这些关键区域,布置了6个高位钻孔,这些钻孔布局合理位置精准,能最大限度地抽取邻近层和采空区中的瓦斯。充分考虑瓦斯运移规律,将钻孔层位选定在砂质泥岩与细砂岩的交界面,距离煤层顶板约30~65 m,以确保钻孔能够最大限度地抽取瓦斯。通过这一高位大直径定向钻孔技术,成功实现了对煤层顶板瓦斯的长期、稳定抽采,使采空区瓦斯涌出量得到了有效控制。在测试结果中,钻场内1#~6#孔平均抽采混量为9.76 m3/min,单孔的最大抽采混量为33.992 m3/min,得出高位大直径定向钻孔在瓦斯抽采过程中具有显著效果,对于提升煤矿瓦斯治理水平具有重要意义。随着煤化工行业的快速发展,提升其技术水平具有重要作用,因此需加强这方面研究,才可以更好地实际应用。

薄煤层工作面切顶卸压留巷技术研究

为提高矿井煤炭资源回收率并提高煤炭开采效率,结合3503综采工作面开采煤层厚度薄的特点,针对性提出切顶卸压留巷技术方案,并进行工程应用。在回风巷切顶卸压留巷期间,通过锚索超前补强可提高留巷段顶板强度,为留巷工作开展创造良好条件;使用DCA-45型切缝机提高切顶钻孔施工质量,通过聚能爆破方式提高切缝效果;在滞后支护段采用密集单体进行补强,可减少采动压力、采空区顶板垮落等对留巷段围岩变形影响。现场应用后,留巷段顶板在滞后采面30 m范围内下沉量增加速度较快,随着与采面距离增加留巷段下沉量逐渐趋稳,最终下沉量控制在165 mm以内。

同忻矿8311工作面坚硬顶板破断特征与矿压显现规律研究

特厚煤层综放开采是煤炭资源开发中的一项重要课题,其开采过程中存在着一系列的难题,其中最为突出的是顶板破断引发强矿压显现。文章基于同忻矿特厚煤层8311工作面,在关键层理论基础上,采用公式计算的方法得出了坚硬顶板的初次和周期来压步距;采用离散单元软件UDEC模拟煤层开挖,得出了坚硬顶板断裂步距与结构。通过现场矿压监测,分析了坚硬顶板的来压步距、周期来压动载强度和“大小周期”现象,探究了坚硬顶板破断与矿压规律之间的关系,为特厚煤层的安全高效开采提供了有力的理论支撑。