水—岩耦合作用下采场底板综合分区特征及其稳定性研究
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摘要
地下水与岩体之间发生机械的、物理的或化学的相互作用,使岩体和地下水的性质或状态不断发生变化,这种相互作用也不断地改变着作用双方的力学状态和力学特性。采矿活动破坏了原始地应力的平衡,采场围岩应力发生重新分布,在新的应力场作用下,采场围岩产生变形和破坏；同时承压水的劈裂、挤入及软化溶蚀作用使岩体劈裂扩展、剪切变形和位移,增加岩体中结构面的孔隙度和连通性,从而增强了岩体的渗透性能,易于形成突水通道,如与采场工作面连通,则发生底板突水。本文以采场底板应力场、位移场演化规律为研究对象,采用理论分析、数值模拟、实验室试验及现场实测等综合研究方法,对底板应力分区、变形分区、破坏分区特征及其相互对应关系进行系统分析,揭示水-岩耦合作用下,不同面长、水压、采厚等条件下底板破坏深度及其稳定性,具体完成内容如下：
(1)根据弹性力学相关理论,将煤层底板视为均质且各向同性的空间半无限体,简化上边界受力形式为weibull分布,得出底板空间应力解析式；
(2)基于水-岩耦合作用原理,对有无水压作用下工作面底板端头位置、中间位置受力情况进行对比分析,结果表明：在承压水含水层附近,有水压作用时底板应力高于无水压作用；端头处高出的应力呈现两次波峰变化形态,中间位置呈现一次波峰变化形态；
(3)水-岩耦合数值分析表明：随含水层水压及面长的增加,含水层中间位置最大主应力呈增大趋势,向两侧逐渐降低；受水压作用,底板垂直应力已降低区域又重新升高,充分卸载范围随水压的增加而减小；孔隙水压力分布规律表明：工作面两端头为突水集中区域,如果煤层开采后剪应力造成两端头处剪切破坏带裂隙发育充分,则容易发生突水事故；
(4)基于采场围岩应力分布规律,得出采场围岩应力特征分区；①载荷缓慢升高区；②端头应力降低区；③拱形应力集中区；④端头关键承载区；
(5)承压水上煤层开采相似模拟试验得出采场底板在采动应力及水压的作用下裂隙产生、发育演化规律。开采初期,在水平应力作用下岩层超出自身抗拉强度出现竖向张裂隙；采动中期,各岩层承载能力增加进而表现出整体抗弯性增强,产生竖向裂隙的同时出现一定量的层间裂隙；采动后期,含水层附近各岩层受水压及其上覆有一定承载能力岩层钳制作用,岩层整体性进一步增强,各岩层不同的挠曲变形造成层间相互错动,形成顺层裂隙；
(6)数字图像相关法对底板移动变形规律进行分析表明：停采线前方35m～40m受超前压力影响,底板出现相对受压状态,其位移值低于未受采动影响区0.2m～0.4m；停采线向采空侧后退5m～10m,底板受压变形达到最大,该范围称之为压缩区；再向后30m～40m时,底板位移恢复到未受采动影响水平,称为压缩臌胀过渡区;离层臌胀区位于压缩膨胀过渡区后65m～75m,该区内产生底臌变形峰值点,最大底臌量为1.2m左右,其中峰值点位于开采范围中间位置后10m～20m左右；从峰值点向后一直延伸到切眼后煤体内10m～20m,底臌变形下降,该区域称为压实稳定区；
(7)模型模拟过程中分析了水压从4.5MPa以0.5MPa压差值进行卸载,水压卸载到1.5MPa时,底臌量最高点下降幅度0.7m,由此表明：就采动应力及底板水压综合作用而言,底板岩层的变形以水压作用为主；
(8)按位移特征整个底板空间分为两个明显区域,采空区对应底板空间大部分范围为底臌区,中部受矸石压实作用效果明显,底臌量有所下降或底臌趋势减缓,形成底臌变形减缓区；随与底板距离增加,下部各岩层承载能力的增强,底臌区内受矸石压实作用影响程度减弱,底臌形态从采场两端向中部逐渐增高,为底臌稳定发展区；含水层上边界到底板不受采动影响深度为底臌变形削弱区,该区域上边界由于受水力下压作用,使得底臌量比非承压水上开采进一步降低,总体亦呈中间高两端低的形态；
(9)相拟材料模拟过程中对底板破坏深度采用并行电法进行实测,结果表明：顶板岩层破坏最显著区域为顶板上方48m,而底板岩层破坏最显著区域为底板下方19m。针对面长120m、160m、200m,水压1MPa、3MPa、5MPa,采厚1.0m、2.0m、3.0m、4.0mm、5.0m等开采条件,进行数值分析采场底板破坏形态及破坏深度,得出适合于面长100m以上,关于以上三因素的底板破坏深度拟合公式：y=57.1ln((?))+0.09M2+0.0644ep-127.727
(10)对采场底板竖向空间从塑性破坏角度进行分带,分别是：充分破坏带、潜在导水破坏带、塑性破坏带。在充分破坏带内形成三个区,采场中部①区原拉破坏基础上产生新的剪切破坏,称重复破坏区；两端头的②区受中部压应力作用与①区力学联系降低,进一步加剧了破坏深度及拉应力受力范围,称之为破坏加剧区。在潜在导水破坏带内中部③区三向应力增大,尽管形成塑性破坏,但仍具备承载能力,称损伤破坏区；两侧剪切面到两端头形成④区,从破坏形式看以拉剪破坏为主,与③区相比三向应力增加幅度不大,为潜在突水区域,称为潜在透水破坏区；
(11)针对A组煤底板岩石进行抗压、抗拉常规性试验得出相应强度值。采用MTS-815液压伺服系统对标准岩石试件进行蠕变试验,得出各级应力水平下试件轴向、横向蠕变参数值；
(12)建立了底板粘弹性岩梁力学分析模型,由虚功原理及能量泛函变分条件,得出采动应力及底板水压不同作用时间下,受粘弹性岩梁抗弯能力下降的影响,底板弹性岩梁挠度及拉应力变化趋势。分析了岩梁弹性模量、粘滞系数及水压变化对其形变影响程度,得出：加固底板以提高岩梁力学性质,一方面增加弹性模量,有助于加强底板抗变形能力,另一方面提高了导水破坏区岩石粘滞系数,减缓完整岩梁受力变形强度；疏水降压可有效降低岩梁边界受力,使其稳定性增强；
(13)针对潘北矿A3煤层地质条件及水文地质条件,结合采场底板破坏深度计算公式,对水压4.5MPa,面长100m、120m、160m、200m时底板破坏深度进行计算,得出为满足安全生产的需要所采取的防治措施；
(14)为保证11113工作面安全回采,确定工作面面长120m；分层开采上分层厚2.8m-3.2m,平均3.0m；实施探放水工程使水压降为0.5MPa。针对现有开采技术条件及水文条件,计算突水系数满足规程要求,同时对工作面底板长期稳定性进行分析表明：现有水文地质及开采技术条件下工作面发生滞后突水可能性很小。对工作面进行现场矿压监测,实测表明初次来压步距38m,周期来压步距18m,工作面推进110m,无底板突水及异常矿压显现。
The mechanism of the interaction of physical、chemical or mechanical between groundwater and rock which not only changes the quality and state of rock and groundwater, but also the mechanical state and the mechanical properties of both sides constantly. Mining activities undermine the balance of the original stress and make stope rock stress redistribute, and stope surrounding rock will produce deformation and failure under new stress field. Meanwhile, the effects of press water splitting, squeeze and soften dissolution make rock mass splitting extension, shear deformation and displacement and increase rock structural plane porosity and connectivity, which increase the permeability of rock mass, be easy to form water inrush channel and make floor water inrush if connecting to stope face. By taking the stope floor stress field, displacement field evolution as the research object, the thesis is done to systematically analyze the stress, deformation and failure partition characteristics and the corresponding relationship between the systematic research on the floor, reveals depth of destroyed floor and its stability of different facial length mining thickness, pressure under the Water-Rock coupling, using comprehensive study methods of similarity simulation test, computer numerical simulation, theoretical analysis and example analysis, etc. The detail research contents are as following:
(1)According to the elastic mechanics theory, the floor of coal seam was regarded as a homogeneous and isotropic half-infinite body and its boundary stress was simplified to the form of Weibull distribution, which can obtain arbitrary stress analysis of the floor space.
(2)Based on the principle of the water-rock interaction, the force of end position and intermediate position of working face floor was comparatively analyzed with water pressure or not, and results showed that floor stress with pressure is higher than the one without pressure near confined aquifers, stress in end position presents two peaks change shape and stress in intermediate position presents one peak change shape.
(3) Water-Rock numerical analysis shows that:With the aquifer water pressure and face length increasing, the maximum principal stress in intermediate position of aquifers tended to increase and gradually reduced to the sides; vertical stress of the floor in stress decreasing zone rises again at the action of pressure, fully uninstall range decreases with the increase of pressure; pore water pressure distribution showed that the two ends of working face were concentrated area of water inrush, it is prone to water inrush if shear failure area cracks at the two ends which are caused by shear stress develop fully after coal mining.
(4) Based on the stope surrounding rock stress distribution, partition characteristics of surrounding rock stress was obtained:①Load slowly increased area;②End stress reduced area;③Arched stress concentration area;④End key bearing area.
(5) Fissure production and development evolution law of stope floor under the stress and pressure were obtained by similar simulation test of coal mining above aquifer. Rock exceeds its tensile strength and vertical tensile cracks occur under the horizontal stress when early mining; bearing capacity of the rock increases and then its overall bending resistance reinforces in mining metaphase, a certain amount of interlayer fissure occur at the same time in vertical cracks occurring; after the mining activities, pressure and overlying rock which has a certain capacity clamps the rock near aquifer and integrity of rock was further enhanced, bedding fissure forms because of the deflection caused by the different rock layers mutual dislocation.
(6) The analysis using digital image correlation method to the movement and deformation laws of the floor shows that:The floor appears relative compression deformation by advanced pressure influence at35cm～40cm ahead of the stopping line, the displacement value is2mm～4mm lower than the one in unaffected zones of coal mining; deformation of the floor gets to the maximum by press at5cm～10cm behind the stopping line towards gob in which is called the formation of the compression zone; Floor displacement restored to the level of mining influence unaffected at more than30cm～40cm behind the stopping line in which is called the compression tympanites transition zone, the separation zone is at65cm～75cm behind compression tympanites transition zone where floor heave deformation peak point appears, the maximum of floor heave reaches about12mm of which the peak point is located10cm～20cm behind the middle position; floor heave deformation decreases from the peak point to10cm～20cm behind the cut in coal body in which is called the restoration of compaction zone.
(7) Floor heave value is analyzed by pressure uninstalled in differential pressure to 0.5MPa from4.5MPa in simulation process, the highest point of floor heave falls7mm when pressure is1.5MPa, it is shown that:As far as the comprehensive force of mining stress and pressure is concerned, floor strata deformation dominated by three gray water pressure.
(8) The entire floor space is divided into two obvious zones by displacement characteristic, most of floor space range corresponded with gob was floor heave zone, floor heave value declines or trend slows down in the central by obvious gangue compaction effect where is formed floor heave deformation reduction zone; with increase of distance to the floor, floor heave form was gradually increased to the central from both ends of the stope by carrying capacity of the lower rock enhanced and influence degree of gangue compaction effect weaken in which is called floor heave stability development zone;the area from the upper boundary of aquifer to the depth not affected by mining in the floor is called floor heave deformation weakening zone, in this area, floor heave is further lower than the one mining without confined water by the region boundary due to hydraulic pressure and the overall is also high in the middle and low in both ends of form.
(9)Measure the floor's failure depth using parallel electrical method in similar material simulation process, the results indicate that the most significant region of roof strata destruction is located at48cm above the roof, and the most significant region of floor strata destruction is located at19cm below the floor. For the mining conditions like face length(120m、160m、200m)、pressure(11MPa、3MPa、5MPa) and mining thickness(1.0m、2.0m、3.0m、4.0m、5.0m), floor failure depth fitting formula can be obtained with analyzing mining floor failure modes and failure dept, which is suitable for face length of more than300and about the above three factors. y=57.11n((?))+0.09M2+0.0644eP-127.727
(10) Stope floor damage divided from the failure point which are A--full damage zone; B--potential water damage zone; C--plastic failure zone. Three areas form band in the full destruction, mining new shear failure of the original field central tension failure basis, called repeated failure zone; two ends of the central area is controlled by the compressive stress and the mechanical contact interval decreased, further aggravating the damage depth and pull stress range, called the destruction area. Three dimensional stress increases in the middle of③zone, despite the formation of plastic damage, but still have the capacity which called the damage zone; both sides of shear surface④form the two ends from failure to shear failure, stress increase is not large compared with③zone, as a potential water inrush area, called the potential permeable failure zone.
(11)Have compressive and tensile strength of conventional testing of the rock group A coal seam and obtain the corresponding intensity value. Have creep test of standard with MTS-815hydraulic servo system and get the specimen axial and lateral creep parameters under different stress levels.
(12)The mechanics analysis model of the floor viscoelastic rock beam is established and the variation tendency of the deflection plate intact rock beam and tensile stress is obtained by the principle of virtual work and energy functional variation condition under the influence of the viscoelastic rock beam flexural capacity decreased in different action time with mining stress and pressure. Analyze influence degree of rock beam elastic modulus, coefficient of viscosity and pressure to the deformation and obtain the result that reinforcement floor can improve the mechanical properties of rock beam, on the one hand, modulus increasing is helpful to strengthen the floor deformation resistance, on another hand, the water area of rock viscous coefficient of destruction improving can slow down the intact rock beam deformation and strength; drainage can effectively reduce the rock beam boundary stress and improve the stability.
(13) For the Panbei Mine A3coal geology and hydrogeology conditions, combining with the stope floor failure depth calculation formula, calculating floor failure depth under the conditions as pressure (4.5MPa) and face length (120m、160m、200m), prevention measurements for safety production are obtained.
(14) In order to ensure the safety mining of11113working face, the working face length is decided to120m, the thickness of higher slice is decided to2.8m-3.2m, mean3.0m and pressure is reduced to0.50MPa using water drainage project. In view of the existing mining conditions and hydrological conditions, calculate inrush coefficient to meet requirements of the regulations and analyze long-term stability of face floor. The results shows that it is likely to occur hysteresis water inrush in face under the existing mining hydrological conditions. For pressure monitoring field to the work surface, the field experiment shows that the first weighting interval is38m and the periodic weighting step distance is18m, there are no water inrush and abnormal underground pressure behavior when advance distance reaches100m.
(1)根据弹性力学相关理论,将煤层底板视为均质且各向同性的空间半无限体,简化上边界受力形式为weibull分布,得出底板空间应力解析式；
(2)基于水-岩耦合作用原理,对有无水压作用下工作面底板端头位置、中间位置受力情况进行对比分析,结果表明：在承压水含水层附近,有水压作用时底板应力高于无水压作用；端头处高出的应力呈现两次波峰变化形态,中间位置呈现一次波峰变化形态；
(3)水-岩耦合数值分析表明：随含水层水压及面长的增加,含水层中间位置最大主应力呈增大趋势,向两侧逐渐降低；受水压作用,底板垂直应力已降低区域又重新升高,充分卸载范围随水压的增加而减小；孔隙水压力分布规律表明：工作面两端头为突水集中区域,如果煤层开采后剪应力造成两端头处剪切破坏带裂隙发育充分,则容易发生突水事故；
(4)基于采场围岩应力分布规律,得出采场围岩应力特征分区；①载荷缓慢升高区；②端头应力降低区；③拱形应力集中区；④端头关键承载区；
(5)承压水上煤层开采相似模拟试验得出采场底板在采动应力及水压的作用下裂隙产生、发育演化规律。开采初期,在水平应力作用下岩层超出自身抗拉强度出现竖向张裂隙；采动中期,各岩层承载能力增加进而表现出整体抗弯性增强,产生竖向裂隙的同时出现一定量的层间裂隙；采动后期,含水层附近各岩层受水压及其上覆有一定承载能力岩层钳制作用,岩层整体性进一步增强,各岩层不同的挠曲变形造成层间相互错动,形成顺层裂隙；
(6)数字图像相关法对底板移动变形规律进行分析表明：停采线前方35m～40m受超前压力影响,底板出现相对受压状态,其位移值低于未受采动影响区0.2m～0.4m；停采线向采空侧后退5m～10m,底板受压变形达到最大,该范围称之为压缩区；再向后30m～40m时,底板位移恢复到未受采动影响水平,称为压缩臌胀过渡区;离层臌胀区位于压缩膨胀过渡区后65m～75m,该区内产生底臌变形峰值点,最大底臌量为1.2m左右,其中峰值点位于开采范围中间位置后10m～20m左右；从峰值点向后一直延伸到切眼后煤体内10m～20m,底臌变形下降,该区域称为压实稳定区；
(7)模型模拟过程中分析了水压从4.5MPa以0.5MPa压差值进行卸载,水压卸载到1.5MPa时,底臌量最高点下降幅度0.7m,由此表明：就采动应力及底板水压综合作用而言,底板岩层的变形以水压作用为主；
(8)按位移特征整个底板空间分为两个明显区域,采空区对应底板空间大部分范围为底臌区,中部受矸石压实作用效果明显,底臌量有所下降或底臌趋势减缓,形成底臌变形减缓区；随与底板距离增加,下部各岩层承载能力的增强,底臌区内受矸石压实作用影响程度减弱,底臌形态从采场两端向中部逐渐增高,为底臌稳定发展区；含水层上边界到底板不受采动影响深度为底臌变形削弱区,该区域上边界由于受水力下压作用,使得底臌量比非承压水上开采进一步降低,总体亦呈中间高两端低的形态；
(9)相拟材料模拟过程中对底板破坏深度采用并行电法进行实测,结果表明：顶板岩层破坏最显著区域为顶板上方48m,而底板岩层破坏最显著区域为底板下方19m。针对面长120m、160m、200m,水压1MPa、3MPa、5MPa,采厚1.0m、2.0m、3.0m、4.0mm、5.0m等开采条件,进行数值分析采场底板破坏形态及破坏深度,得出适合于面长100m以上,关于以上三因素的底板破坏深度拟合公式：y=57.1ln((?))+0.09M2+0.0644ep-127.727
(10)对采场底板竖向空间从塑性破坏角度进行分带,分别是：充分破坏带、潜在导水破坏带、塑性破坏带。在充分破坏带内形成三个区,采场中部①区原拉破坏基础上产生新的剪切破坏,称重复破坏区；两端头的②区受中部压应力作用与①区力学联系降低,进一步加剧了破坏深度及拉应力受力范围,称之为破坏加剧区。在潜在导水破坏带内中部③区三向应力增大,尽管形成塑性破坏,但仍具备承载能力,称损伤破坏区；两侧剪切面到两端头形成④区,从破坏形式看以拉剪破坏为主,与③区相比三向应力增加幅度不大,为潜在突水区域,称为潜在透水破坏区；
(11)针对A组煤底板岩石进行抗压、抗拉常规性试验得出相应强度值。采用MTS-815液压伺服系统对标准岩石试件进行蠕变试验,得出各级应力水平下试件轴向、横向蠕变参数值；
(12)建立了底板粘弹性岩梁力学分析模型,由虚功原理及能量泛函变分条件,得出采动应力及底板水压不同作用时间下,受粘弹性岩梁抗弯能力下降的影响,底板弹性岩梁挠度及拉应力变化趋势。分析了岩梁弹性模量、粘滞系数及水压变化对其形变影响程度,得出：加固底板以提高岩梁力学性质,一方面增加弹性模量,有助于加强底板抗变形能力,另一方面提高了导水破坏区岩石粘滞系数,减缓完整岩梁受力变形强度；疏水降压可有效降低岩梁边界受力,使其稳定性增强；
(13)针对潘北矿A3煤层地质条件及水文地质条件,结合采场底板破坏深度计算公式,对水压4.5MPa,面长100m、120m、160m、200m时底板破坏深度进行计算,得出为满足安全生产的需要所采取的防治措施；
(14)为保证11113工作面安全回采,确定工作面面长120m；分层开采上分层厚2.8m-3.2m,平均3.0m；实施探放水工程使水压降为0.5MPa。针对现有开采技术条件及水文条件,计算突水系数满足规程要求,同时对工作面底板长期稳定性进行分析表明：现有水文地质及开采技术条件下工作面发生滞后突水可能性很小。对工作面进行现场矿压监测,实测表明初次来压步距38m,周期来压步距18m,工作面推进110m,无底板突水及异常矿压显现。
The mechanism of the interaction of physical、chemical or mechanical between groundwater and rock which not only changes the quality and state of rock and groundwater, but also the mechanical state and the mechanical properties of both sides constantly. Mining activities undermine the balance of the original stress and make stope rock stress redistribute, and stope surrounding rock will produce deformation and failure under new stress field. Meanwhile, the effects of press water splitting, squeeze and soften dissolution make rock mass splitting extension, shear deformation and displacement and increase rock structural plane porosity and connectivity, which increase the permeability of rock mass, be easy to form water inrush channel and make floor water inrush if connecting to stope face. By taking the stope floor stress field, displacement field evolution as the research object, the thesis is done to systematically analyze the stress, deformation and failure partition characteristics and the corresponding relationship between the systematic research on the floor, reveals depth of destroyed floor and its stability of different facial length mining thickness, pressure under the Water-Rock coupling, using comprehensive study methods of similarity simulation test, computer numerical simulation, theoretical analysis and example analysis, etc. The detail research contents are as following:
(1)According to the elastic mechanics theory, the floor of coal seam was regarded as a homogeneous and isotropic half-infinite body and its boundary stress was simplified to the form of Weibull distribution, which can obtain arbitrary stress analysis of the floor space.
(2)Based on the principle of the water-rock interaction, the force of end position and intermediate position of working face floor was comparatively analyzed with water pressure or not, and results showed that floor stress with pressure is higher than the one without pressure near confined aquifers, stress in end position presents two peaks change shape and stress in intermediate position presents one peak change shape.
(3) Water-Rock numerical analysis shows that:With the aquifer water pressure and face length increasing, the maximum principal stress in intermediate position of aquifers tended to increase and gradually reduced to the sides; vertical stress of the floor in stress decreasing zone rises again at the action of pressure, fully uninstall range decreases with the increase of pressure; pore water pressure distribution showed that the two ends of working face were concentrated area of water inrush, it is prone to water inrush if shear failure area cracks at the two ends which are caused by shear stress develop fully after coal mining.
(4) Based on the stope surrounding rock stress distribution, partition characteristics of surrounding rock stress was obtained:①Load slowly increased area;②End stress reduced area;③Arched stress concentration area;④End key bearing area.
(5) Fissure production and development evolution law of stope floor under the stress and pressure were obtained by similar simulation test of coal mining above aquifer. Rock exceeds its tensile strength and vertical tensile cracks occur under the horizontal stress when early mining; bearing capacity of the rock increases and then its overall bending resistance reinforces in mining metaphase, a certain amount of interlayer fissure occur at the same time in vertical cracks occurring; after the mining activities, pressure and overlying rock which has a certain capacity clamps the rock near aquifer and integrity of rock was further enhanced, bedding fissure forms because of the deflection caused by the different rock layers mutual dislocation.
(6) The analysis using digital image correlation method to the movement and deformation laws of the floor shows that:The floor appears relative compression deformation by advanced pressure influence at35cm～40cm ahead of the stopping line, the displacement value is2mm～4mm lower than the one in unaffected zones of coal mining; deformation of the floor gets to the maximum by press at5cm～10cm behind the stopping line towards gob in which is called the formation of the compression zone; Floor displacement restored to the level of mining influence unaffected at more than30cm～40cm behind the stopping line in which is called the compression tympanites transition zone, the separation zone is at65cm～75cm behind compression tympanites transition zone where floor heave deformation peak point appears, the maximum of floor heave reaches about12mm of which the peak point is located10cm～20cm behind the middle position; floor heave deformation decreases from the peak point to10cm～20cm behind the cut in coal body in which is called the restoration of compaction zone.
(7) Floor heave value is analyzed by pressure uninstalled in differential pressure to 0.5MPa from4.5MPa in simulation process, the highest point of floor heave falls7mm when pressure is1.5MPa, it is shown that:As far as the comprehensive force of mining stress and pressure is concerned, floor strata deformation dominated by three gray water pressure.
(8) The entire floor space is divided into two obvious zones by displacement characteristic, most of floor space range corresponded with gob was floor heave zone, floor heave value declines or trend slows down in the central by obvious gangue compaction effect where is formed floor heave deformation reduction zone; with increase of distance to the floor, floor heave form was gradually increased to the central from both ends of the stope by carrying capacity of the lower rock enhanced and influence degree of gangue compaction effect weaken in which is called floor heave stability development zone;the area from the upper boundary of aquifer to the depth not affected by mining in the floor is called floor heave deformation weakening zone, in this area, floor heave is further lower than the one mining without confined water by the region boundary due to hydraulic pressure and the overall is also high in the middle and low in both ends of form.
(9)Measure the floor's failure depth using parallel electrical method in similar material simulation process, the results indicate that the most significant region of roof strata destruction is located at48cm above the roof, and the most significant region of floor strata destruction is located at19cm below the floor. For the mining conditions like face length(120m、160m、200m)、pressure(11MPa、3MPa、5MPa) and mining thickness(1.0m、2.0m、3.0m、4.0m、5.0m), floor failure depth fitting formula can be obtained with analyzing mining floor failure modes and failure dept, which is suitable for face length of more than300and about the above three factors. y=57.11n((?))+0.09M2+0.0644eP-127.727
(10) Stope floor damage divided from the failure point which are A--full damage zone; B--potential water damage zone; C--plastic failure zone. Three areas form band in the full destruction, mining new shear failure of the original field central tension failure basis, called repeated failure zone; two ends of the central area is controlled by the compressive stress and the mechanical contact interval decreased, further aggravating the damage depth and pull stress range, called the destruction area. Three dimensional stress increases in the middle of③zone, despite the formation of plastic damage, but still have the capacity which called the damage zone; both sides of shear surface④form the two ends from failure to shear failure, stress increase is not large compared with③zone, as a potential water inrush area, called the potential permeable failure zone.
(11)Have compressive and tensile strength of conventional testing of the rock group A coal seam and obtain the corresponding intensity value. Have creep test of standard with MTS-815hydraulic servo system and get the specimen axial and lateral creep parameters under different stress levels.
(12)The mechanics analysis model of the floor viscoelastic rock beam is established and the variation tendency of the deflection plate intact rock beam and tensile stress is obtained by the principle of virtual work and energy functional variation condition under the influence of the viscoelastic rock beam flexural capacity decreased in different action time with mining stress and pressure. Analyze influence degree of rock beam elastic modulus, coefficient of viscosity and pressure to the deformation and obtain the result that reinforcement floor can improve the mechanical properties of rock beam, on the one hand, modulus increasing is helpful to strengthen the floor deformation resistance, on another hand, the water area of rock viscous coefficient of destruction improving can slow down the intact rock beam deformation and strength; drainage can effectively reduce the rock beam boundary stress and improve the stability.
(13) For the Panbei Mine A3coal geology and hydrogeology conditions, combining with the stope floor failure depth calculation formula, calculating floor failure depth under the conditions as pressure (4.5MPa) and face length (120m、160m、200m), prevention measurements for safety production are obtained.
(14) In order to ensure the safety mining of11113working face, the working face length is decided to120m, the thickness of higher slice is decided to2.8m-3.2m, mean3.0m and pressure is reduced to0.50MPa using water drainage project. In view of the existing mining conditions and hydrological conditions, calculate inrush coefficient to meet requirements of the regulations and analyze long-term stability of face floor. The results shows that it is likely to occur hysteresis water inrush in face under the existing mining hydrological conditions. For pressure monitoring field to the work surface, the field experiment shows that the first weighting interval is38m and the periodic weighting step distance is18m, there are no water inrush and abnormal underground pressure behavior when advance distance reaches100m.
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