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1.
Zhongchang Wang Xinping Guo Chuan Wang 《Geotechnical and Geological Engineering》2018,36(6):3399-3409
To master the laws of strong strata behavior of Tashan coal mine under Carboniferous coal mining process, the laws of strong strata behavior in 8107 working face was measured and analyzed. It was shown that the average initial weighting step of 8107 working face was 59.4 m. The average periodic weighting step of main roof was 16.2 m. The maximum working resistance during periodic weighting was 14,711.1 kN. The maximum working resistance during non-periodic weighting was 11,339.9 kN. The average dynamic load factor K during periodic weighting was 1.31. The stress of coal column on the side of the goaf could be divided into four zones (stress stabilization zone, stress slow-increasing zone, significant—increasing stress zone, stress reduction zone) along the strike of 8107 working face. There was a peak of lateral support pressure along the trend of 8107 working face. And the peak position was biased to the side of return airway roadway. With the increase of the distance from the down-side of return airway, the pressure peak of the inner coal body along the strike of 8107 the working face increased and the peak position decreased from the coal wall. The peak stress of coal column tended to be close to the up-side of return airway. And the distance from the down-side of return airway for the peak of inner coal was larger than that for the peak of coal pillar. The peak position of abutment pressure of hard roof was in the range of 10–25 m in front of 8107 working face under full mechanized mining extra thickness coal seam conditions. The relative stress concentration coefficient of k was 1.3–6.5. The range of 10–25 m from the front of the working face to coal wall was stress reduction zone. And the influence range of abutment pressure was about 80 m. It was of great significance to the control and practice of the surrounding rock of the stope for the mining of the hard extra-thick coal seam. 相似文献
2.
Yushun Yang Dongming Zhang Xin Bai Han Yang 《Geotechnical and Geological Engineering》2018,36(4):2087-2095
In this paper, based on the field test of No.S3012 working face of Shan Mushu Coal Mine in Sichuan Coal Group, monitoring the abutment pressure and gas drainage flow during the mining process, studying the change law of the abutment pressure and gas drainage flow of the coal seam, and using the numerical simulation method research on the evolution of abutment pressure and displacement of coal seam during the mining process. The results shown that: with the advance of coal mining face, the abutment pressure of coal seam can be divided into stress decreasing area, stress increasing area and original stress area, and the stress state of coal seam and the pore, crack structure and permeability of coal body are obviously changed. With the advance of the mining face, the abutment pressure in front and back of the coal mining face is the moving abutment pressure, and the coal mining face to be in the pressure relief area, the front abutment pressure peak value deep into the coal body 5–10 m, the influence scope reaches the front coal mining face to 90–100 m, this area is the stress increasing area. And the evolution law of the roof displacement of goaf is similar to the elliptical with the axial ratio changes, when the ratio is close to 1, the roof subsidence affected area is similar to the shape of “O”. 相似文献
3.
Some villages and bridges are located on the ground surface of the working district no. 7 in the Wanglou Coal Mine. If longwall mining is adopted, the maximum deformation of the ground surface will exceed the safety value. Strip mining is employed for the working district no. 7 which is widely used to reduce surface subsidence and the consequent damage of buildings on the ground surface. To ensure the safety of coal pillars and improve the recovery coefficient, theoretical analysis and numerical simulation (FLAC 3D) were adopted to determine the coal pillar and mining widths and to discuss the coal pillar stress distribution and surface subsidence for different mining scenarios. The results revealed that the width of coal pillars should be larger than 162 m, and the optimized mining width varies from 150 to 260 m. As the coal seam is exploited, vertical stress is mainly applied on the coal pillar, inducing stress changes on its ribs. The coefficient of mining-induced stress varies from 2.02 to 2.62 for different mining scenarios. The maximum surface subsidence and horizontal movement increase as the mining width increases. However, when the mining width increases to a certain value, increasing the pillar width cannot significantly decrease the maximum subsidence. To ensure the surface subsidence less than 500 mm, the mining width should not be larger than 200 m. Considering the recovery coefficient and safety of the coal pillar, a pillar width of 165 m is suggested. 相似文献
4.
Borehole stress monitoring analysis on advanced abutment pressure induced by Longwall Mining 总被引：1，自引：0，他引：1
Nong Zhang Nianchao Zhang Changliang Han Deyu Qian Fei Xue 《Arabian Journal of Geosciences》2014,7(2):457-463
The paper presented the research on the dynamic advanced abutment stress induced by longwall mining with borehole stress meters on mining side coal mass. Twenty vibrating wire borehole stress meters were installed into the extracting coal mass wall of a first mining roadway of 910 m depth in Zhuji Coal Mine, China, and were used to monitor dynamic changes in vertical and horizontal stresses. Three months of continuous monitoring and further analysis showed that the impacting distance of advanced abutment stress induced by mining in the strike of the working face along its central axis was the farthest, greater than 250 m (the face length is 220 m); it gradually decreased in the radial direction of the face from its central axis outward; the pressure peak was located within 24 m in the front of the mining coal wall; non-synchronous caving of the layered mudstone roof at the stope occurred. Comparison between vertical and horizontal stress increments indicated that the horizontal stress was much smaller than the vertical stress in the coal mass of mining side, while the latter’s magnitude determined the drastic degree of mine pressure manifestation. The study has been applied to determine the advanced support length of the working face and further provide a reliable basis to forecast such dynamic disasters as rock burst, coal and gas outburst, etc., as well as to design the asymmetric supports on both sides of a gateway. 相似文献
5.
Qi Wang Hongke Gao Bei Jiang Shucai Li Manchao He Dechao Wang Wei Lu Qian Qin Song Gao Hengchang Yu 《Arabian Journal of Geosciences》2017,10(21):466
Driving roadway along a goaf is commonly adopted for mining face of thick seam in a deep mine. Determining a reasonable width of coal pillar is a key scientific problem for driving roadway along a goaf in a deep mine. The paper took a roadway driven along a goaf at Zhaolou coal mine which is a typical kilometer-deep mine in China as engineering background. Field monitoring, model test, and numerical experiment are conducted. Stress and displacement evolution mechanism are analyzed with different pillar widths. The test results show that with the increase of coal pillar width, the peak stress value at the coal pillar working slope and integrated coal beside the roadway increases firstly and then tends to be stable, its position is transferred to the side of the roadway, and the deformation of coal pillar decreases gradually during roadway excavation. The coal pillar deformation and roadway vertical displacement increased as the coal pillar width increases under high abutment pressure. In order to reduce the waste of non-renewable resources and meet the requirements of bearing capacity and stability of coal pillars, a method is proposed for setting a reasonable width of coal pillars and the specific width of coal pillars is designed and applied in engineering practices based on the above research. All the tests are significant in the study of driving roadway along a goaf in a deep mine. 相似文献
6.
近断层煤层开采时,防隔水煤岩柱的正确留设是防止断层突水事故的重要措施。山东济宁三号煤矿123上01(南)工作面西邻的F8断层落差较大,可能会沟通底部奥陶系灰岩(简称奥灰)含水层,为确保工作面安全回采,需要进行F8断层防隔水煤岩柱尺寸计算。考虑煤岩层真实产状、断层面与煤层产状的空间关系和底板承压水威胁,对《煤矿防治水细则》中断层导水条件下的防隔水煤岩柱尺寸计算方式进行改进,推导出水位到安全防隔水煤岩柱宽度(Ha)在断层面垂足处的水头高度计算公式及改进后的断层防隔水煤岩柱尺寸计算公式。对比计算结果发现:《煤矿防治水细则》原有计算公式在将煤岩层理想化为水平岩层,且取偏小的煤层底板水压值计算的情况下,计算出的F8断层煤岩柱尺寸为112 m,改进公式的计算结果为128.5 m。因改进公式考虑了实际煤岩层产状及与断层面空间关系,且水压取值位置准确,计算结果更为精确,将为矿井安全回采提供更为科学的参考依据。 相似文献
7.
侧向支承压力分布、资源回收率以及煤柱和巷道的稳定性是大采高综放面区段煤柱宽度留设要兼顾的因素,为了确定大采高综放面区段煤柱宽度,以某矿8103面为工程背景,首先,采用理论计算和现场应力监测等方法确定大采高综放工作面倾向支承压力分布规律,得出应力降低区宽度约为8 m,原岩应力区为巷帮侧28 m外。其次,采用工程类比方法确定大采高综放工作面巷帮外侧煤体严重破裂区宽度约为4 m。最后,采用FLAC3D数值软件分析了下区段工作面回采时窄煤柱（6、8 m）和宽煤柱（28、30 m）的应力场、位移场及塑性区特征,获得不同煤柱宽度时巷道和煤柱力学特征。研究表明：当煤柱宽度6 m和8 m时,在采动支承压力下煤柱几乎无承载能力,且巷道变形量较大;当煤柱宽度28 m和30 m时,在采动支承压力下煤柱中央仍有一定的弹性核,煤柱保持稳定且巷道变形量较小。综合考虑资源回收、巷道稳定性、次生灾害控制等因素,确定大采高综放工作面区段煤柱宽度为28 m。 相似文献
8.
C型采场支承压力分布特征的数值模拟研究 总被引：2，自引：0，他引：2
冲击地压的发生与支承压力的分布有重要关系。为研究C型采场支承压力的动态变化规律,采用FLAC3D软件对孤岛工作面推进过程中煤体垂直应力场进行了数值模拟。通过对工作面推进过程中煤体支承压力的平面分布特征、走向支承压力和倾斜支承压力的动态演化特征进行分析,得到以下结论：① 煤体中垂直应力分布呈“C”形;② 孤岛工作面超前支承压力影响距离为正常工作面的3～5倍;③ 双工作面“见方”时,支承压力峰值达到最大值。工程实例验证了结论的可靠性,其结果可为现场冲击地压预测和防治提供依据。 相似文献
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11.
浅埋近水平煤层采动地面变形规律研究 总被引：9，自引：2，他引：7
以神府矿区大柳塔煤矿１２０３综采工作面为例,探讨了浅埋煤层开采过程中,地面下沉变形孤过程及有关参数,为地面建筑物的安全煤柱留设提供了依据。『 相似文献
12.
煤矿突水溃砂灾害的发生与煤层上履含水层性质、岩性特征及破坏程度等诸多因素有关。通过研究己15煤层顶板基岩与第四系底部的含、隔水性能及顶板覆岩岩性组合特征,计算出一次全部开采3．6m煤层时,其导水裂缝带最小发育高度为38．46m,最大为47．95m,确定了采煤活动导致的上覆顶板含水层发生水力联系的范围,认为己15—13030工作面煤层开采时发生顶板突水的可能性不大;计算一次全部开采3．6m煤层时,防砂安全垂高最大为23．5m．防塌安全煤岩柱最大垂高为12．5m,结合煤层顶板基岩及第四系底部岩层的水文地质特征,认为工作面回采时顶板溃砂的可能性也不大。强调在生产过程中,要加强顶板涌水的观测,同时增加现有排水系统的排水能力,从而为工作面安全回采提供支持。 相似文献
13.
浅埋深、薄基岩、厚松散含水层下煤层综合机械化开采防治水技术应用 总被引：1，自引：0，他引：1
神东煤炭公司大柳塔矿12404工作面煤层埋深浅，上覆基岩薄、松散含水层厚，且地表有河流横过(工作面长度240m，推进度3900m，其中薄基岩富水段150m)，为了减少工作面搬家次数提高生产效益，采取了地面抽排水、井下泄放水及地面注浆固沙等有效的防治水措施，工作面安全通过了母河沟，该防治水方法的应用改变了以往类似情况进行留煤柱跳采的方法，取得了浅埋深薄基岩厚含水层下煤层综合机械化安全开采经验，比停产跳采节省搬家费用2400万元，获得了显著的经济效益。 相似文献
14.
通过对唐口煤矿地应力、3上煤层及顶板岩层冲击地压测试结果分析,认为3上煤层属强冲击倾向性煤层,3上煤层顶板属弱冲击倾向性岩层;在采深1000m条件下,随着地应力的增大,煤、岩层的冲击倾向性将会增大。因3上煤层为易碎煤,厚度较大,顶板弹性能易突然全部释放,形成冲击地压;3上煤顶板主要为中砂岩、细砂岩及泥岩,质地坚硬,在煤层开采过程中,煤壁附近出易现高应力集中带,在顶板中聚集的弹性能在自重力和采掘干扰下会突然释放,形成冲击地压。在生产过程中采取钻屑法、沿采煤工作面轨道顺槽安装顶板离层报警系统、合理开拓避免应力集中和叠加、对煤层进行注水,降低煤体弹性和强度、提高支护结构的承载能力等一系列措施,较好地预防了冲击地压的发生。 相似文献
15.
Numerical Investigation of the Dynamic Mechanical State of a Coal Pillar During Longwall Mining Panel Extraction 总被引：2，自引：1，他引：1
Hongwei Wang Yaodong Jiang Yixin Zhao Jie Zhu Shuai Liu 《Rock Mechanics and Rock Engineering》2013,46(5):1211-1221
This study presents a numerical investigation on the dynamic mechanical state of a coal pillar and the assessment of the coal bump risk during extraction using the longwall mining method. The present research indicates that there is an intact core, even when the peak pillar strength has been exceeded under uniaxial compression. This central portion of the coal pillar plays a significant role in its loading capacity. In this study, the intact core of the coal pillar is defined as an elastic core. Based on the geological conditions of a typical longwall panel from the Tangshan coal mine in the City of Tangshan, China, a numerical fast Lagrangian analysis of continua in three dimensions (FLAC3D) model was created to understand the relationship between the volume of the elastic core in a coal pillar and the vertical stress, which is considered to be an important precursor to the development of a coal bump. The numerical results suggest that, the wider the coal pillar, the greater the volume of the elastic core. Therefore, a coal pillar with large width may form a large elastic core as the panel is mined, and the vertical stress is expected to be greater in magnitude. Because of the high stresses and the associated stored elastic energy, the risk of coal bumps in a coal pillar with large width is greater than for a coal pillar with small width. The results of the model also predict that the peak abutment stress occurs near the intersection between the mining face and the roadways at a distance of 7.5 m from the mining face. It is revealed that the bump-prone zones around the longwall panel are within 7–10 m ahead of the mining face and near the edge of the roadway during panel extraction. 相似文献
16.
多煤层开采条件下煤层覆岩破坏具有独特的特征，影响矿井生产布置。以陕北某矿为例，以该矿地质采矿条件为基础，采用相似材料模拟实验与数值模拟相结合的方法，通过建立模拟模型，开展了双煤层开采对覆岩的破坏影响研究。结果显示:留设不同宽度的煤柱，采用相似材料模拟和数值模拟2种方法得到的煤层覆岩垮落带高度、裂隙带高度都基本一致；在双煤层开采时，留设的煤柱宽度越大，两个煤层的叠置区域就越小，煤层开采对覆岩的破坏程度就越小。在工作面布置时，建议增大两个煤层的开采距离，并尽量增加煤柱宽度，以减缓覆岩移动破坏范围和破坏程度。研究成果为类似双煤层开采工作面的设计及覆岩破坏控制提供技术支撑。 相似文献
17.
断层对顶板稳定性影响相似模拟试验研究 总被引：13，自引：1，他引：12
通过相似模拟试验方法分析了不同倾向高角度正断层, 在采动影响下顶板岩体变形破坏和矿压分布规律。结果表明, 在采动影响下断层“活化”,断层带及其影响范围内的岩体破碎, 表现为周期断裂步距小, 冒落带高, 尤其是断层下盘, 顶板稳定性差; 当工作面开采到离断层面22.5~ 30 m时, 直到断层位置的前方煤体中支承压力增大, 煤体被压碎, 且随着距断层面距离的缩小, 支承压力的峰值位置向工作面前方转移; 通过断层后, 顶板岩体中支承压力减小, 比无断层存在的情况要低。 相似文献
18.
为了提供提高煤层开采上限的根据，开展了新集井田东翼11-2煤层防水煤柱厚度、岩性组合特征及含水性的研究，井进行了推复体夹片地层下井田东翼11-2煤层的开采实验，在26m防水煤柱下实现1111200工作面安全回采，取得较好的经济效益。 相似文献
19.
采动岩体变形与渗透特性是工作面突水防治研究的基本问题。采用理论分析和数值模拟计算方法,分析了岩石变形-渗透特征及其三维定量关系,研究了煤炭开采中回采工作面围岩应变场、渗透系数场分布及其控制因素。研究结果表明,采动岩体渗透性变化主要取决于应变状态及应变增量,且随着垂直于裂隙的张应变的增加而增加;工作面后方垮落带和煤壁边缘的剪碎带产生剧烈的采动拉伸变形,渗透系数较采前显著增大,而支承压力区和整体移动带岩层产生较大的采动压缩变形,渗透系数较采前明显减小;工作面围岩垂向渗透系数较水平渗透系数增加的幅度及增加区的范围小,但对水体下采煤工作面涌水起主导作用。加快工作面推进速度、减小工作面斜长和采高能降低采场围岩渗透系数增加幅度,并将渗透系数场变化范围局限在采区附近,能有效减小采动对原始煤岩层渗透性的影响。 相似文献
20.
以神府矿区大柳塔煤矿1203综采工作面为例，建立了覆岩移动的有限元地质模型，探讨了浅埋近水平煤层采动条件下的岩移和塌陷的岩体应力位移变化，为合理设计开采方案、减少塌陷范围、保护生态环境提供科学依据。 相似文献