Stress state and caving danger of the roof in bolt supporting roadway

The start point of this text is the bottleneck problem of bolt supporting coal entrythat is security problem of bolt supporting roof,we divide one entry into some sections withdifferent stress,simulate stress field of wall rock and rockbolt solidified at different sectionsused umbrella disperse soft UDEC(universal distinct element code),we educe that thestress level of wallrock and bolt solidified is higher in roof fall risk section,and roof rockboltload can reflect this rule clearly,that offer an important guideline in monitoring entry rooffall risk.

Roof pre-blasting to prevent support crushing and water inrush accidents

Support crushing and water inrush when mining under an unconsolidated confined aquifer in the Qidong Coal Mine was prevented by roof pre-blasting. The mechanism and applicable conditions for this method have been studied. The results show that when an overburden structure that may cause support crushing and a water inrush accident exists the weakening of the primary key stratum, which thereby reduces its weighting step, roof pre-blasting is both feasible and effective. If the position of the primary key stratum can be moved upward to exceed 10 times the mining height the possibility of support crushing and water inrush disaster caused by key stratum compound breakage will be lowered. The overburden structure of the number 7121 working face was considered during the design of a technical proposal involving roof pre-blasting. After comprehensively analyzing the applicability of roof pre-blasting the resulting design prevented support crushing and water inrush disasters from happening at the number 7121 working face and laid a solid foundation for mining safely.

Quantification of ventilation enhancement using the Eye CAN roof support

Convergence of roof and floor in underground mine openings is a common occurrence. This convergence not only adversely affects the ability of workers, equipment and supplies to travel through the mine, it also reduces the effectiveness of the mine ventilation system, which is essential for the dilution of methane gas and airborne respirable dust. While installing secondary standing supports to control floor and roof convergence, such supports, by nature, partially obstruct a portion of the airway. These added obstructions inhibit the ability of the ventilation system to operate as efficiently as it could by increasing the resistance in and reducing the cross-sectional area of the airway. This study introduces and demonstrates the benefits of The Eye CAN^(TM) standing roof support, which controls floor and roof convergence and is less obstructive to air flow than conventional wooden cribs. Laboratory findings show that the normal resistance of a supported lined airway is reduced by using this new product from Burrell Mining Products, Inc., while providing the same roof support characteristics of an established product—The CANò. Load vs. displacement curves generated from laboratory tests demonstrated that this new product behaves with the same roof support characteristics as others in The CAN product family. Ventilation data gathered from a simulated mine entry was then used for computational fluid dynamics(CFD) modeling.The CFD analysis showed an improvement with The Eye CAN vs. other accepted forms of standing roof support. This proof-of-concept study suggests that, when using this new product made by Burrell Mining Products, Inc., not only will the convergence from the roof and floor be controlled, but airway resistance will also be reduced.

Influence of canopy ratio of powered roof support on longwall working stability——A case study

The case study describes longwall coal seam A in a hard coal mine,where longwall coal face stability loss and periodic roof fall occurrences had been registered.The authors have attempted to explain the situation based on in-situ measurements and observations of the longwall working as well as numerical simulation.The calculations included several parameters,such as powered roof support geometry in the form of the canopy ratio,which is a factor that influences load distribution along the canopy.Numerical simulations were realized based on a rock mass model representing realistic mining and geological conditions at a depth of 600 m below surface for coal seam A.Numerical model assumptions are described,while the obtained results were compared with the in-situ measurements.The conclusions drawn from this work can complement engineering knowledge utilized at the stage of powered roof support construction and selection in order to improve both personnel safety and longwall working stability,and to achieve better extraction.

Mode of overlying rock roofing structure in large mining height coal face and analysis of support resistance

The mining space of large mining height coal face is large,the range of movement and caving of rock strata is large and the stability of supports at coal face is low and damage rate of supports is high,which significantly affects the safe and efficient production of coal mines.By similar simulation experiment and theoretical analysis,the mode of fractured roofing structure of large mining height coal face and the method of determination of reasonable support resistance of the support was evaluated.Analysis shows that the structural mode of "combined cantilever beam – non-hinged roofing – hinged roofing" of the large mining height coal face appears at the roofing of large mining height coal face.The supporting factor of caved gangue at the gob is introduced,the calculating equations of the fractured step distance of roofing were derived and conventional calculating method of caved height of roofing was corrected and the method of determination of the length and height of each structural area of the roofing was provided.With reference to the excavating conditions at Jinhuagong coal mine in Datong minefield,the dimensions of structural areas of the roofing of the coal face were determined and analyzed,and reasonable support resistance of the height coal face was acquired.By selecting Model ZZ13000/28/60 support and with procedures of advanced pre-cracking blasting,the safe production of large mining height coal face was assured.

Horizontal roof gap of backfill hydraulic support

For the backfill hydraulic support as the key equipment for achieving integration of backfilling and coal mining simultaneously in the practical process, its characteristics will directly influence the backfill body's compression ratio. Horizontal roof gap, as a key parameter of backfilling characteristics, may impact the backfilling effect from the aspects of control of roof subsidence in advance, support stress, backfilling process and the support design. Firstly, the reason why horizontal roof gap exists was analyzed and its definition, causes and connotation were introduced, then adopting the Pro/E 3D simulation software, three typical 3D entity models of backfill hydraulic supports were built, based on the influence of horizontal roof gap on backfilling effect, and influence rules of four factors, i.e. support height, suspension height, suspension angle and tamping angle, were emphatically analyzed on horizontal roof gap. The results indicate that, the four factors all have significant impacts on horizontal roof gap, but show differences in influence trend and degree, showing negative linear correlation, positive linear correlation, positive semi-parabolic correlation and negative semi-parabolic correlation, respectively. Four legs type is the most adaptive to the four factors, while six legs(II) type has the poorest adaptability, and the horizontal roof gap is small under large support height, small suspension height, small suspension angle and large tamping angle situation. By means of optimizing structure components and their positional relation and suspension height of backfill scrape conveyor in the process of support design and through controlling working face deployment, roof subsidence in advance, mining height and backfilling during engineering application, the horizontal roof gap is optimized. The research results can be served as theoretical basis for support design and guidance for backfill support to have better performance in backfilling.

Roof structure theory and support resistance determination of longwall face in shallow seam

This paper presents the structure models founded in shallow seam, the roof asymmetry arch with three articulations in roof first weighting and the step voussoir beam in roof periodic weighting. These structure models are differ from classic theory, it establishes the new roof control theory of instability structure roof, especially in shallow seam. Based on the new roof structure theory, the support working state of "given sliding load" is put forward, and the factor of load transmitting is introduced to determine the load on roof structure. Therefore, the proper and accurate calculating methods of support resistance are established. Based on this, the dynamic structure theory in shallow seam could be predicted.

Bolt Supporting of Large-Span Soft Rock Roadway in Shaqu Colliery

The instability of trapezoidal Ⅰ-steel support is analysed for the compound roof of main coal seam in Shaqu Colliery, and the mechanism of bolt supporting is studied. A scheme of bolt supporting has been given and put into practice, remarkable technical and economic benefits have been got.

Maximal loads acting on legs of powered roof support unit in longwalls with bumping hazards

In the article the results of measurements of the resultant force in the legs of a powered roof support unit, caused by a dynamic interaction of the rock mass, are discussed. The measurements have been taken in the longwalls mined with a roof fall, characterized by the highest degree of bumping hazard. It has been stated that the maximal force in the legs F m, recorded during a dynamic interaction of the rock mass, is proportional to the initial static force in the legs F st,p . Therefore a need for a careful selection of the initial load of the powered roof support, according to the local mining and geological conditions, results from such a statement. Setting the legs with the supporting load exceeding the indispensable value for keeping the direct roof solids in balance, deteriorating the operational parameters of a longwall system also has a disadvantageous influence on the value of the force in the legs and the rate of its increase, caused by a dynamic interaction of the rock mass. A correct selection of the initial load causes a decrease in the intensity of a dynamic interaction of the rock mass on powered roof supports, which also has an advantageous influence on their life. Simultaneously with the measurements of the resultant force in the legs, the vertical acceleration of the canopy was also recorded. It has enabled to prove that the external dynamic forces may act on the unit both from the roof as well as from the floor. The changes of the force in the legs caused by dynamic phenomena intrinsically created in the roof and changes of the force in the legs caused by blasting explosives in the roof of the working, have been analyzed separately. It has been stated that an increase in the loads of legs, caused by intrinsic phenomena is significantly higher than a force increase in the legs caused by blasting. It means that powered roof supports, to be operated in the workings, where the bumping hazard occurs, will also transmit the loads acting on a unit during blasting. The majority of recorded force changes in the legs has been caused by a dynamic interaction of the roof. They are characterized by a load increase coefficient K d, satisfying the inequality 1 06<K d =F m /F st,p <1 24. A much smaller number of cases, when the external load acted on the bases, was recorded. Individual, recorded results of measurements indicate that changes of the force in the legs, caused by external loads of this type, run more intensively due to roof loads (1 08< K d<1 80),particularly in these cases when the near the roof layer of the seam is under mining. A determination of more precise relations among the changes of forces in the legs, caused by a dynamic interaction of the floor and the bases and the mining and geological conditions requires a performance of additional underground tests.

巷帮煤体整体滑脱型冲击地压锚杆防冲支护原理及工程实践

冲击地压是目前严重影响煤炭安全有效开采的灾害之一,研究锚杆防冲支护原理和技术对防治巷道冲击地压灾害具有重要意义和价值。通过对巷帮煤体整体冲入型冲击地压发生的地质条件以及破坏特征进行总结分析,认为坚硬顶板与坚硬煤层是此类型冲击地压的重要地质特征,而巷帮煤体整体滑脱是其主要冲击破坏特征。在此基础上,以巷帮滑脱煤体为研究对象,建立了顶板-巷道-底板复合结构体力学模型,建立了巷帮煤体发生水平滑移的极限平衡方程,并对各个参数进行分析。结果表明:由于顶板反弹使巷帮煤体竖直方向压力降低,巷帮煤体被构造应力推入巷道发生冲击地压。基于该发生机制模型,认为目前的锚杆支护设计体系在防治巷帮煤体冲入型冲击地压存在不足,并基于其发生和破坏特征,建立了针对巷帮煤体整体滑脱型冲击地压的锚杆防冲支护设计原则,即将顶和底帮锚杆锚固端分别穿层打入稳定的顶底板内,并使用长锚索取代中部帮锚杆,提供锚杆支护的防冲作用。基于新建立的锚杆防冲支护设计方法,以大屯矿区孔庄煤矿7305工作面防冲支护为工程背景,在宽煤柱段巷帮锚杆支护采取了防冲设计,帮顶、底部锚杆及补强锚索均锚固于顶底板内部,能够有效吸收煤体滑动动能,提高安全性。

十堰某体育馆106m跨内凹弦支穹顶结构设计关键问题研究

十堰某体育馆主体结构采用框架-剪力墙结构体系,上部106m大跨内凹形屋盖采用弦支穹顶结构体系。为了满足建筑简洁性及结构合理性需求,对屋盖进行了四角锥网架、弦支穹顶及网架-弦支穹顶组合的不同结构形式对比,最终选用结构轻巧、经济性好的弦支穹顶结构体系。结合弦支穹顶结构自身的受力特点以及其对下部结构的不利影响,对弦支穹顶结构上下弦布置、扭转刚度、支座约束条件、撑杆高度等进行了优化。对弦支穹顶断索情况下的响应、不同工况下的支座位移以及拉索低应力状态下的刚度退化进行了验算。结果表明,优化后的屋盖结构满足结构安全、适用性的要求,同时也有较高的冗余度。

基于力学模型构建的留巷切顶高度确定与围岩控制技术

以顺和煤矿2401运输巷道沿空留巷为工程背景,分析巷道围岩结构特征,基于巷道局部空间结构稳定性,分别构建了沿空留巷未切顶与切顶力学结构模型,并以巷道不同切顶高度进行物理相似模拟试验。结果表明:巷旁采空区切落的矸石增加对关键块的支撑力,同时弱化关键块对直接顶悬壁端部挤压,巷旁支护阻力减少35.78%;随着切顶高度增加,巷道顶板采空区侧端部悬臂由F型缓慢过渡到大I型,同时大保护结构具有向上平移趋势,相比于未切顶1巷,4 cm切顶2巷、8 cm切顶3巷与16 cm切顶4巷叠加应力峰值分别下降9.38%,28.13%,25.00%。结合巷道顶板岩性,最终确定切顶高度为8.2 m,留巷段采用三列单体液压支柱作巷旁支护,长短锚索超前补强,巷道围岩稳定,较好满足使用要求。

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。

大采高综采工作面矿压显现及顶板活动特征研究

为了得到唐山某煤矿综采工作面的矿压显现及顶板活动特征,基于关键层理论、“悬臂梁-砌体梁”力学模型,通过现场观测、理论分析、数值模拟的方法对9煤层矿压显现特征、顶板结构、支架工作阻力开展了深入研究,建立了覆岩结构力学模型,得到了工作面支架工作阻力的合理计算式。研究结果表明:工作面来压步距和来压大小均有一定的规律性,呈现一大一小的周期来压现象,来压期间中部测区支架荷载最大,两边偏小;通过直接顶关键层判别方法,确定了第2层岩层和第7层岩层分别为亚关键层1和亚关键层2。亚关键层1破断后以“悬臂梁”结构形态垮落进入采空区,亚关键层2破断后形成了稳定的“砌体梁”结构,此时亚关键层1与亚关键层2共同组成了“悬臂梁-砌体梁”结构;第2层岩层和第7层岩层组成的双关键层结构的破断运移规律导致了工作面呈现一大一小的周期来压现象。当双关键层同步破断时,支架工作阻力为13 688 kN。

基于能量法的工作面端面冒顶机理及“支架-围岩”耦合关系试验研究

为研究综采工作面顶板稳定性及其影响因素,结合能量法和“顶板-支架-煤壁”系统模型,建立了端面顶板稳定性力学模型,利用分布式支架顶梁压力监测系统和数字图像监测技术,开展了端面冒顶相似模拟试验,分析了端面冒顶影响因素、顶板破坏形态和支架顶梁压力分布演化特征。研究表明:根据端面顶板稳定性力学模型,距煤壁1 m范围内直接顶稳定性系数小于0,该区域为端面冒顶高发区;距离煤壁越远,顶板垂直位移和水平位移也随之增大;提高支架工作阻力能有效减小顶板下沉量;顶板黏聚力和内摩擦角越大,顶板稳定性越好;相似模型试验中直接顶依次经历了端面冒落、顶板破断、顶板破碎等阶段,液压支架顶梁压力表现为中部>前部>后部;当顶板完整时,模型支架初撑力和支护阻力充足,“支架-围岩”耦合关系良好;当顶板破碎导致支架位态不佳时,模型支架降阻或偏载现象明显,容易造成顶板裂隙发育、端面冒顶、支架压架等事故,“支架-围岩”耦合关系恶化;利用数字图像监测技术获得了端面冒顶阶段直接顶最大剪应变集中于端面顶板附近。研究显示支架与围岩相互作用关系与端面顶板稳定性互相影响,维持良好的“支架-围岩”耦合关系有助于提高工作面顶板稳定性。

新型异形曲面钢⁃木组合结构临时支撑卸载方案优化与监测

温州未来城区体验馆屋面结构采用异形曲面钢木组合,立面采用纤细的两端铰接的钢立柱,钢立柱同时作为玻璃幕墙的支承结构,地上抗侧力墙体采用钢骨混凝土墙体,地下室为混凝土结构。充分利用了木结构和钢结构的优势,最大程度减少结构的用钢量。但由于结构跨度大且形状复杂,施工阶段存在安装偏差,临时支撑拆卸需要较高的同步性及稳定性,需要基于实际起拱坐标和荷载状态,对其卸载过程进行计算分析和施工监测以确保结构安全。基于实际施工情况,运用MIDAS/GEN对该组合结构的卸载施工过程进行了数值模拟计算,通过已有的初步方案明确了各施工步中的杆件内力以及节点位移,并在此基础上提出了分步分级的优化方案,同时对其在卸载施工过程中的位移响应进行了动态监测。计算结果表明:通过对52个千斤顶进行分步卸载,支撑点位移与构件应力成倍增加,采取分级卸载策略,逐步减小荷载,成功实现了对结构的精准控制,避免了内部过度变形和应力集中,确保了结构稳定性。监测点的实际值与理论值误差在合理范围内,验证了模拟分析的准确性,为工程实践提供了可靠支持。

基于三线性应变软化模型的综采工作面煤壁片帮深度解析预测

为预测综采工作面煤壁片帮深度,以煤壁为研究对象,煤壁煤体视作三线性应变软化材料;将直接顶和基本顶视作宏观顶板,建立了宏观顶板与煤层、支架相互作用的承载力学模型,结合边界条件推导了综采工作面煤壁片帮深度解析计算公式,利用该公式研究上湾煤矿12401工作面煤壁片帮情况,分析影响煤壁片帮的因素及影响规律,结果表明:①推导获得的煤壁片帮深度计算公式能够较准确计算综采工作面煤壁片帮深度,能够定量分析顶板特性和支架参数等对煤壁片帮深度的影响。②增加支架初撑力能够降低煤壁片帮深度;增加支架刚度能够降低煤壁片帮深度,但效果较差;增加支架护帮板支护强度,煤壁片帮深度近似线性减小。③顶板厚度和抗拉强度增加,煤壁片帮深度增长,但增长速率随顶板厚度和抗拉强度增加逐渐趋缓。这主要由于顶板厚度和抗拉强度增加,顶板破断步距增大,向采空区内旋转下沉角度增大所致。④综采采高对煤壁片帮深度有很大影响,综采采高增大,煤壁片帮深度近似线性增长。⑤煤体单轴抗压强度增大,煤壁片帮深度减小,顶底板与煤层界面之间的摩擦因数增大,煤壁片帮深度减小。⑥顶板动载系数增大,煤壁片帮深度近似线性增大。

顶板支护条件下的切顶工程优化研究及应用

以金厂河ZnV3矿体阶段空场嗣后充填采场顶部切顶条柱稳定性为研究对象,通过分析采场在不同阶段应力状态的分布规律,结合顶板破碎围岩锚固支护的现状,为提升条柱的稳定性,采用理论分析和数值模拟相结合的方法进行研究分析。研究结果表明:切顶联络道两侧交替布置4 m宽凿岩硐室、3 m宽条柱的优化方案,能有效分担应力集中作用对条柱的影响,极大提高了条柱的稳定性;优化方案改变了顶板支护区域,进一步提高了空区后顶板的稳定性。根据研究成果开展了工业试验,验证了优化方案的可行性,对安全生产具有重要意义,可在矿山进行应用。切顶工程的优化,在提高条柱稳定性的情况下,进一步提高了顶板的稳定性,可为类似矿山提供借鉴。