Localization of mining-induced horizontal fractures along rock layer interfaces in overburden: field measurements and prediction

The locations of mining-induced horizontal fractures along rock interfaces in the overburden of Donetsk Coal Basin were identified using an original experimental device. The device traps methane from horizontal fracture zone (100–fold coal seam thickness) over an active longwall mining excavation. Presence or absence of horizontal fractures along rock layer interfaces is correlated with physical characteristics of the overburden, such as thickness, uniaxial compressive strength of overburden rock layers, location of rock layer interfaces and thickness of extracted coal seams. As a result, a combined criterion based on these physical characteristics is proposed to predict the presence of overburden horizontal fracturing in coal mine operations.     

A RBFNN-based method for the prediction of the developed height of a water-conductive fractured zone for fully mechanized mining with sublevel caving

The movement and failure of overlying strata induced by underground coal mining cause “three zones,” including the caving zone, the water-conductive fractured zone, and the sagging zone from the bottom up. For knowledge about the height of the water-conductive fractured zone, there has been no empirical or theoretical formulae for thick coal seam using fully mechanized longwall mining with sublevel caving. This paper presents a methodology of determining the height of the water-conductive fractured zone based on the radial basis function neural networks (RBFNN) model in MATLAB software. Before modeling, the relationship between the height of the water-conductive fractured zone and mining thickness, the lithologic character of the overburden and its composite structures, and workface parameters was studied. After that, 32 and 7 measured data were used as training and testing samples, respectively. It has been found that the average relative error is 6% and the maximum relative error is 10% for 7 test samples by comparing actual results with predicted results. The model was applied to the no. 31503 workface in the Gaozhuang coal mine for safety evaluation. The predicted value is 59.6 m, and the measured value is 55.9 m. The RBF-based model shows much better performance than empirical formulae in the Regulations for Coal Mining and Coal Pillar Design under Buildings, Water-bodies, Railways and Main Shafts for the prediction of the height of the water-conductive fractured zone for fully mechanized mining with sublevel caving.     

A study of surface subsidence and coal pillar safety for strip mining in a deep mine

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.     

多煤层开采覆岩破断过程的模型试验与数值模拟

首先分析了研究区煤矿山西组煤层剖面和水文地质特征,针对多煤层联合开采的特点和覆岩的工程地质特征,采用工程地质力学模型实验、数值模拟计算相结合的综合研究方法,分析了多煤层开采的采动影响及岩层动态断裂机理,得出了有关岩层移动参数和多层煤同采时应力分布状态,计算得到了多煤层开采垮落带与导水裂隙带的发育高度分别为32m和81.5m,导水裂隙带高度影响范围已经达到风化带,未形成切冒,局部穿透粘土层。     

Measurement of the Strong Pressure Appearance Laws with Hard Roof and Full Mechanized Mining Extra Thick Coal Seam

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.     

Grout injection into bed separation to control surface subsidence during longwall mining under villages: case study of Liudian coal mine,China

Surface subsidence can cause many environmental problems and hazards (including loss of land area and damage to buildings), and such hazards are particularly serious in coal mining districts. Injecting grout into the bed separation in the overburden has been proposed as an effective control measure against surface subsidence during longwall mining. However, no field trials of this technique have been implemented in mines under villages in China, and thus, its ability to control subsidence in such areas has yet to be demonstrated. In this study, field trials using this technique were carried out during longwall mining under villages in the Liudian coal mine, China. The maximum surface subsidence observed after the extraction was only 0.298 m, which accounts for 10 % of the mining height and is 79 % less than the predicted subsidence. Moreover, no damage occurred to the village buildings either during or after extraction and these buildings remain stable. Thus, this study represents the first successful attempt to control surface subsidence under villages in China using grout injection during longwall mining.     

废弃采空区裂隙带高度预测模型及应用

煤炭开采引起覆岩破断及地表下沉,覆岩及地表运移规律可反映裂隙带高度的动态演化过程。因地表下沉滞后于煤炭开采,对于废弃采空区,长期压实作用导致裂隙带高度较采动期间有所降低。基于地表点下沉速度的阶段特征将裂隙带高度的演化过程分为2个阶段,第1阶段裂隙带发育对应岩层破断逐步向上传递的过程,第2阶段裂隙带高度降低对应离层及裂隙闭合、断裂岩层受压后变形回弹及破碎岩体自然压实的过程。着眼于压实作用对裂隙带高度的影响,根据煤层采厚、垮落带和裂隙带岩层变形量及地表下沉值之间的定量关系,建立了第2阶段裂隙带高度预测模型,并结合太平煤矿实测结果进行验证,采用控制变量法分析了单一因素影响下废弃采空区裂隙带高度的演化特征。结果表明：废弃采空区裂隙带高度受控于垮落带块体强度、垮落带初始碎胀系数、采动期间裂隙带高度最大值及对应的垮落带高度、煤层埋深、地表最终下沉量等因素,太平煤矿采后15 a的裂隙带高度实测值11.36～13.00 m与理论预测值12.75 m吻合度较高,模型的可靠性得到验证。最后,应用此预测模型对武安煤矿（关停矿井）2002－2003年采空区裂隙带高度开展理论计算,结合地空瞬变电磁探测确定了地面瓦斯抽采钻孔理想的终孔位置并成功开展了地面钻孔瓦斯抽采试验。     

黄河流域陕北煤炭开采区厚砂岩对覆岩采动裂隙发育的影响及采煤保水建议

煤炭开采活动导致的煤层顶板覆岩地质条件变化及采动裂隙发育是损害地下关键含水层的直接原因,也是造成矿区生态环境退化的根源。煤层顶板覆岩结构中发育的厚砂岩作为一种典型的地质条件,其对覆岩采动裂隙的发育规律具有重要的影响。为此,在分析研究区主采煤层赋存地质条件及其分布规律的基础上,选择陕北煤炭开采区曹家滩煤矿主采2?2煤层顶板覆岩为地质原型,采用FLAC3D数值模拟平台模拟分析了厚砂岩不同厚度和位置对覆岩采动裂隙发育形态和发育高度的影响,并以此提出了相应的“采煤保水”建议。结果表明:研究区2?2煤层顶板覆岩中厚砂岩平均厚度25 m,距2?2煤层平均间距76 m;厚砂岩距煤层30 m时,覆岩采动裂隙表现为“矩形—L形—马鞍形”的动态变化特征,距煤层70 m时表现为“L形—倒梯形—马鞍形”变化特征,距煤层大于95 m时全程表现为“马鞍形”特征;覆岩采动裂隙最大发育高度随厚砂岩层位的升高而先减小后增大;厚砂岩厚度H≥30 m、距煤层间距L>95 m,或H≥60 m、L>60 m时,可有效阻挡采动裂隙向上发育贯穿厚砂岩;在充分考虑厚砂岩对覆岩采动裂隙发育规律的影响,选择合适的空间位置和开采阶段进行合理的覆岩减损和保水防治,实现“边采边治、边采边护”的绿色开采模式。该研究成果可为黄河流域中游陕北煤矿区煤炭开采与生态环境保护协调发展提供理论指导。      

深部条带开采覆岩“三带”探测及量化评判

为研究深部条带开采覆岩裂隙发育高度这一采空塌陷区治理中的关键技术问题，综合运用工程地质钻探、钻孔电视与煤田测井3种方法对济宁煤田某煤矿覆岩采动裂隙进行探测与分析。结果表明:缓倾斜（煤层平均倾角6°）、深埋（平均埋深538 m）、采留比1:2（采50 m留100 m）的条带式采动下，覆岩“三带”特征显著，空间整体呈“波浪式”破坏形态，地下采空区无明显空洞；垮落带发育高度为8.45 m，为采高的3.0倍，断裂带发育高度为57.55 m，为采高的20.6倍；覆岩“三带”发育高度判别的8个量化指标中，钻孔电视法比工程地质钻探法、煤田测井法精确度高。探测结果较为准确可靠，为深部条带采空塌陷区治理方案的选择提供依据。      

Fracture Evolution and Accumulation and Dissipation Law of Energy During Ascending Mining

Ascending mining method is a widely used method in China for coal mine safety. To determine whether ascending mining method can eliminate or reduce the outburst risk or not in Xinzyao Mine, China, a numerical model was built by FLAC3D software to investigate the fracture evolution and accumulation and dissipation law of energy during the process of ascending mining. And a dual water blocking system was used to verify the fracture evolution obtained from the numerical model. The results show that FLAC3D model can predict the fracture evolution in overburden with a reasonable error. The height of water-conducting fractured zone of no. 14-3 coal seam is about 82.5 m, and it increases by 27 % after no. 12-2 coal seam is mined. The largest energy accumulated in no. 12-2 coal seam is about 32.5 kJ/m³ when mining it directly. The energy accumulated in no. 12-2 coal seam decreases by more than 40 % when ascending mining method is adopted. Ascending mining method can effectively eliminate the outburst risk of the upper coal seam. The research is helpful for the determination of ascending mining method and the further study of the prevention mechanism of rock burst.     

浅埋煤层过沟开采覆岩裂隙与地表裂缝演化规律数值模拟

浅埋煤层过沟开采在陕北矿区普遍存在,严重威胁着矿井安全生产和生态健康发展。覆岩采动裂隙及地表裂缝的发育是岩体微裂隙的延伸和扩展的结果,为更加系统地研究浅埋煤层过沟开采覆岩微裂隙的演化规律,以陕北安山井田125203工作面为背景,根据现场沟道剖面及钻孔资料,采用Particle Flow Code (PFC)数值模拟平台构建颗粒流数值采煤模型,模拟分析覆岩微裂隙的发育特征、数量变化规律和力链演化特征,揭示微裂隙的发育规律和地表裂缝发育机理。结果表明:浅埋煤层过沟开采过程中,覆岩微裂隙发育表现为“产生–延伸和扩展–聚合成群–贯通成缝”的动态演化过程;根据微裂隙的基本发育特征及分布规律,可将发育全程划分为非连续跳跃式、连续贯通式和横向扩展式3个微裂隙发育阶段;覆岩微裂隙数量随工作面推进距离的增大而增多,非连续跳跃式和连续贯通式发育阶段呈现指数增长特征,分别累计发育微裂隙547和2 867条,覆岩微裂隙逐渐发育至地表;横向扩展式发育阶段呈线性增长特征,累计发育微裂隙11 705条,微裂隙数量随岩层高度的增大而减小;覆岩力链演化过程中,强黏性力链的破坏致使微裂隙发育,局部产生应力集中,强黏性力链拱自下而上逐渐破坏并贯通至地表,导致微裂隙延伸发育至地表形成地裂缝,力链在未贯通岩层的微裂隙两侧和尖端及地表裂缝两侧形成强力链区,微裂隙处形成弱(无)力链区。该研究成果可为陕北矿区浅埋煤层过沟开采覆岩及地表损害防控提供理论指导。      

基于微震监测的深埋煤层顶板导水裂隙带发育特征

导水裂隙带发育高度是矿井水害预测的重要技术参数之一。以彬长矿区文家坡煤矿4103工作面为研究对象,利用井-地联合微震监测技术对顶板导水裂隙带发育特征进行研究。研究结果表明:深埋煤层开采时,微震事件超前工作面回采位置发育,超前影响角最大为35°,最小为28°;断层的存在降低了覆岩稳定性,相较于正常基岩,更易在回采影响下发生应力集中和破坏;断层加大了微震事件发生的超前距,而采空区则使微震事件的高密度区向其所在部位发生偏移,加剧覆岩破坏程度,增大导水裂隙带发育高度;垂向上,4103工作面监测区内的微震事件高密度区域主要集中在高程+400~+520 m,结合微震事件数量和能量分布特征,判定4103工作面垮落带发育高度为50 m,垮采比13.16,导水裂隙带发育高度为117 m,裂采比为30.79。该成果可为彬长矿区类似煤矿深埋煤层顶板导水裂隙带发育高度研究及顶板水防治提供重要依据。移动阅读      

Coal Mine Roadway Stability in Soft Rock: A Case Study

Roadway instability has always been a major concern in deep underground coal mines where the surrounding rock strata and coal seams are weak and the in situ stresses are high. Under the high overburden and tectonic stresses, roadways could collapse or experience excessive deformation, which not only endangers mining personnel but could also reduce the functionality of the roadway and halt production. This paper describes a case study on the stability of roadways in an underground coal mine in Shanxi Province, China. The mine was using a longwall method to extract coal at a depth of approximately 350 m. Both the coal seam and surrounding rock strata were extremely weak and vulnerable to weathering. Large roadway deformation and severe roadway instabilities had been experienced in the past, hence, an investigation of the roadway failure mechanism and new support designs were needed. This study started with an in situ stress measurement programme to determine the stress orientation and magnitude in the mine. It was found that the major horizontal stress was more than twice the vertical stress in the East–West direction, perpendicular to the gateroads of the longwall panel. The high horizontal stresses and low strength of coal and surrounding rock strata were the main causes of roadway instabilities. Detailed numerical modeling was conducted to evaluate the roadway stability and deformation under different roof support scenarios. Based on the modeling results, a new roadway support design was proposed, which included an optimal cable/bolt arrangement, full length grouting, and high pre-tensioning of bolts and cables. It was expected the new design could reduce the roadway deformation by 50 %. A field experiment using the new support design was carried out by the mine in a 100 m long roadway section. Detailed extensometry and stress monitorings were conducted in the experimental roadway section as well as sections using the old support design. The experimental section produced a much better roadway profile than the previous roadway sections. The monitoring data indicated that the roadway deformation in the experimental section was at least 40–50 % less than the previous sections. This case study demonstrated that through careful investigation and optimal support design, roadway stability in soft rock conditions can be significantly improved.     

Experimental Study on Displacement Field of Strata Overlying Goaf with Sloping Coal Seam

To study the displacement caused by underground mining, the displacement field in the strata overlying the mine was simulated in nine similar-material models with coal seam dip angles of 0°–80°. Digital close-range photogrammetry was employed to observe the displacement in these models and produce displacement vector diagrams of observation points set on the model surfaces. The movement boundary of the rock mass was extracted and determined. According to the displacement vectors’ direction, the displacement field within the movement boundary was divided into five zones: two zones with displacement vectors pointing towards the goaf center, two zones with vectors pointing towards the coal pillar, and one zone with displacement vectors pointing vertically downward. A symmetry index was defined to analyze the symmetry of the displacement field, and the results show that as the coal seam dip angle increased from 0° to 80°, the displacement field in the strata changed according to the following pattern: symmetric → asymmetric → nearly symmetric, with a transition angle of about 45°. The percentage of the area of the above-defined five zones relative to the displacement field’s total area also changed, showing a regular pattern. This study also revealed that the movement boundary inside the overlaying rock layers is not a straight line, but an S-shaped curve. These findings can enhance our understanding of the internal mechanism of ground subsidence and contribute to more effective prediction of the deformation occurring inside a given rock mass.     

Detecting the footprint of a longwall mine panel claimed to infringe on a permit boundary at the Soma–Darkale coalfield (Manisa, Turkey) using surface fractures and microgravity measurements

A coal mine panel claimed to infringe on a permit boundary, or to create an hazard after subsidence has been the subject of lawsuits. We study at the Soma–Darkale coalfield (Manisa, Turkey), the footprint of a lignite coal mine panel at a depth of about 150–200 m by mapping all of the surface fractures we could observe, and by developing a post-subsidence density model that we verified through gravity measurements with positive Bouguer anomaly. With the analysis of the fracture map and the gravity data, we were able to identify the footprint of a mine panel from the effect of the anomalous mass due to denser overburden material filling up the space after the extraction of less-dense lignite. Whereas, using empirical methods like the so-called “limit line approach” to evaluate the extent of the area where mining can have subsidence-induced surface fracturing, one could not recognize for certain infringement of permit boundaries. The orientations of the fractures we mapped at the ground surface, the Bouguer gravity map, and a test borehole indicated the presence and dimensions of a coal panel in dispute. The presented approach based on gravity method and fracture observations may be an example to help settle conflicts related to the position of the longwall mine panel.     

特厚煤层大采高综放工作面煤壁片帮机制分析

针对特厚煤层大采高工作面,将超前支承压力作用的煤体分为破裂区、塑性区和弹性区。考虑支架护帮阻力等因素,建立了支承压力作用的煤体变形破坏及煤壁片帮分析力学模型。通过摩尔-库仑准则和非关联弹塑性分析,推导出了煤壁水平位移量及破裂区和塑性区半径解析表达式,分析了影响煤壁片帮的主要因素。理论结果与现场实测结果误差仅为15.6%,验证了模型的正确性。研究表明,随支承压力集中系数的增大、机采高度的增加以及支架护帮阻力的减小,煤壁水平位移量明显增大,煤壁片帮程度严重。工程实践表明,提高支架工作阻力,减缓煤壁处压力,控制合理采高以及增大支架护帮阻力,可有效缓解大采高综放工作面煤壁片帮程度。     

双系煤层开采相互影响下的覆岩运动与破坏规律分析

以大同矿区同忻煤矿为例,利用关键层理论、物理探测技术和数值模拟,分析双系煤层开采形成的覆岩运动与破坏规律。研究结果表明：（1）石炭二叠系煤层工作面推进至193 m时,覆岩裂隙带发育高度达到最大为174.7 m;（2）物理探测得到3-5^#煤层开采引起的覆岩垮裂带高度约为180 m,综合理论计算与现场实测,确定覆岩裂隙带发育高度与采厚之比为11.7-12.0;（3）侏罗系煤层开采引起的底板破坏及双系间覆岩软弱夹层的缺失为双系煤层采空区联通进一步提供了条件;（4）侏罗系煤层的开采影响石炭二叠系煤层工作面超前支承压力的分布规律,当石炭系煤层工作面位于侏罗系煤层遗留煤柱下方时,超前支撑压力较无采空区时增大12.4%;（5）研究成果揭示了双系开采相互影响关系,为采取安全开采措施、减弱双系开采相互影响程度提供了科学依据。     

榆神矿区保水采煤的工程地质背景

论述了榆神矿区煤层上覆红土、风化带基岩、未风化基岩的工程地质特征,认为红土层具有较好的隔水性能,风化带基岩中粘土矿物的含量增加,隔水性能明显增强。2-2煤层上覆基岩为中硬岩石,而且有较多的软弱岩石夹层,煤层开采的冒裂带发育高度相对较小,影响不到第四系萨拉乌苏组含水层底部,具有较好的保水采煤工程地质条件。      

测井方法在观测采空区上覆岩层导水裂隙带发育高度中的应用

新安矿区主采煤层上部至小浪底库区水体之间地层主要以砂岩、泥岩和砂质泥岩互层为主,厚度90~210m。为研究该矿开采状态下上覆岩层破坏程度和导水裂隙发育高度,在井下3个不同采厚的工作面上布置了7个地面钻孔,采用声波扫描成像测井及其它测井参数对采前、采后上覆岩层进行动态观测。以K3钻孔为例,介绍了超声波扫描成像测井的原理及辨别井孔中裂缝发育程度的方法。K3孔在第一次测量时,井深217～218m处明显存在裂隙,但在第二次测量后该裂隙呈现出闭合反应,证实煤层开采放顶后上部岩石下沉使得原有裂隙闭合;同时发现自194m以上因煤层采空放顶后发育有高度为69.10m的裂隙带。根据新安煤矿K2、K3、K5、K6、K7号孔超声波扫描成像及其它常规测井的地质解释成果可见,该矿导水裂隙带的发育高度随着推进距离和采厚增大而增大,但新安矿区软硬互层上覆岩层的地质构造可有效抑制导水裂隙带的发育程度。     

Physical simulation research on evolution laws of clay aquifuge stability during slice mining

The aquifuge stability is the key to study the impacts of coal mining on the aquifer. Based on the geological conditions of a mine in Yili of Xinjiang, China, this paper has studied the stability evolution laws of clay aquifuge during extremely thick coal seam mining by similar material simulation experiment in the laboratory. For the water-swelling and expansion property of clay aquifuge, the reasonable proportion of the similar material is firstly determined by taking the uniaxial compressive strength and the permeability coefficient as core indexes. Then, the overlying strata movement coupled solid–liquid physical model is established. In addition, the aquifuge deformation, the water level changes of the aquifer, and the height of fracture zone in overburden are analyzed during the slice mining. The research results indicate that the clay aquifuge will gradually occur instability failure during the mining of the working face, and the aquifuge stability has the threshold effects. When the ratio of the vertical displacement of the aquifuge to the thickness is D_v/T?≤?58.0%, the ratio of the horizontal displacement to the thickness is D_h/T?≤?17.0%, and the height of fractured zone in overburden is below the aquifuge, the mining-induced fractures may be closed and the aquifuge stability could be maintained. If D_v/T?≥?75.0%, D_h/T?≥?23.9%, and the height of fractured zone in overburden is within the aquifuge, the fractures will develop and connect the aquifuge and the stability failure of the aquifuge will occur, which has a direct correlation with the mining height.