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.

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.

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.

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.

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.

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。

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。

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。

Preventing roof fall fatalities during pillar recovery:A ground control success story

For decades, pillar recovery accounted for a quarter of all roof fall fatalities in underground coal mines.Studies showed that a miner on a pillar recovery section was at least three times more likely to be killed by a roof fall than other coal miners. Since 2007, however, there has been just one fatal roof fall on a pillar line. This paper describes the process that resulted in this historic achievement. It covers both the key research findings and the ways in which those insights, beginning in the early 2000 s, were implemented in mining practice. One key finding was that safe pillar recovery requires both global and local stability.Global stability is addressed primarily through proper pillar design, and became a major focus after the2007 Crandall Canyon mine disaster. But the most significant improvements resulted from detailed studies that showed that local stability, defined as roof control in the immediate work area, could be achieved with three interventions:(1) leaving an engineered final stump, rather than extracting the entire pillar,(2) enhancing roof bolt support, particularly in intersections, and(3) increasing the use of mobile roof supports(MRS). A final component was an emphasis on better management of pillar recovery operations.This included a focus on worker positioning, as well as on the pillar and lift sequences, MRS operations,and hazard identification. As retreat mines have incorporated these elements into their roof control plans,it has become clear that pillar recovery is not ‘‘inherently unsafe." The paper concludes with a discussion of the challenges that remain, including the problems of rib falls and coal bursts.

顶升式梯形钢管混凝土支架支护技术研究及应用

白芨沟煤矿0102303工作面运输巷道位于上分层采空区下方,顶板为金属网假顶,原支护受上分层采空区扰动影响,变形严重。为此设计顶升式梯形钢管混凝土支架,通过注浆顶升试验测试各型号支架顶升性能,试验结果表明:Φ194 mm×10 mm钢管注浆效果较好,接顶力达到3.93 kN,满足支架顶升和巷道假顶初撑要求。为增加钢管梁抗弯强度,在空钢管内设置抗弯圆钢,计算得支护反力为2.3 MPa,较原支护方式提高了5.4倍。同时,为解决巷道内各类支架安装困难,设计了一种举升式支架安装机,举升力可达6.8 kN。顶升式梯形钢管混凝土支架注浆30 d后,支架荷载基本区域稳定,安装半年内支架荷载最大值为0.8 MPa,现场监测顶底板移进和两帮收敛均小于50 mm,支护效果良好。

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。

韩城矿区破碎围岩巷道注浆锚索加固机理研究

以陕西韩城桑树坪矿3315运输巷道为研究背景,通过对注浆锚索破碎围岩巷道加固作用的研究,揭示了注浆锚索的加固机理,提出了软弱煤层巷道支护设计的原则,并进一步优化支护方案,通过采用高强度锚索对围岩进行加固,提升围岩的整体稳定性,达到有效防止围岩变形和损伤的效果。根据监测结果得出,加固后巷道顶板下沉量由原支护方式的帮部收敛量平均为32.7 mm,顶板的平均下沉量为28.6 mm,围岩的平均变形量仅为原支护方案的15.6%,顶板变形稳定后,巷道顶板离层量较小,围岩稳定后平均深基点位移仅为原支护方案的55%。说明优化后的支护方案能够很好控制巷道的围岩变形,有效控制了锚固区范围内围岩的离层,减小了顶板弯曲沉降,増强了顶板的稳定性,该方案可有效控制围岩变形。研究为软弱煤层巷道的支护设计提供了重要的理论依据和工程经验,并为软弱煤层巷道支护提供借鉴和指导。

龙凤煤矿切顶卸压成巷围岩控制研究

为了提高煤炭采出率,基于龙凤煤矿生产条件,通过理论分析、数值模拟、现场监测等方法研究切顶卸压沿空留巷控制对策,提出“恒阻大变形锚索+单体和十字梁+可缩性U型钢”的联合支护方式。通过对巷道位移量、液压支架压力、巷道漏风量、沿空留巷甲烷浓度监测数据的分析,沿空留巷滞后工作面约130 m时围岩开始趋于稳定,巷道漏风率在2.14%~3.80%内,风量满足使用要求,沿空留巷回风流中的瓦斯浓度在留巷期间均低于1%,没有甲烷超限现象发生。经过现场观测,巷道顶板预裂切缝率高,液压支架压力明显降低,顶板卸压效果明显,取得了良好的留巷效果。

瓯海奥体中心体育馆100m跨无环索弦支网壳结构设计

瓯海奥体中心由体育场、体育馆、游泳馆及体育配套用房组成,其中体育馆主体结构采用钢筋混凝土框架结构,屋盖采用了无环索弦支网壳结构体系,为国内外首例该结构体系工程。设计中针对屋盖结构进行了屈曲分析及抗连续倒塌分析,结果表明结构的整体稳定性较好、抗连续倒塌能力强,进一步证明了结构体系选择的合理性。对拉索索力进行了专项分析,保证在各工况组合下拉索均不破断及退出受拉状态,且刚度退化后仍满足设计要求。

厚煤层大采高综采面顶板破断规律及机理研究

以杭来湾煤矿首个大采高30201工作面为例,通过数值模拟、理论分析和现场实测的方法,分别对工作面开采过程中顶板破断规律进行研究。基本顶以“悬臂梁”、“砌体梁”的结构形式周期发生破断,造成工作面大、小周期来压显现。建立了工作面顶板力学模型,给出来压步距的计算方法。现场实测表明最大步距67.6 m是平均步距34.4 m的1.97倍,说明约2次小周期来压才有1次大周期来压。

大跨度破碎岩体顶板预应力锚固支护机理及效果研究

大跨度破碎岩体顶板采用预应力锚固支护,能最大程度地维持顶板围岩稳定,因此研究预应力锚固对大跨度破碎岩体顶板的支护机理及支护效果,对于保障破碎岩体下大跨度地下工程、露天大断面隧道工程的长期稳定性具有重要意义。对锚固后锚固体强度变化进行力学解析,剖析大跨度破碎岩体顶板预应力锚固支护机理,以云南金厂河多金属矿Ⅳ级破碎岩体大跨度顶板支护难题为工程背景,针对其支护方式建立数值模型进行分析,对FLAC3D软件中结构单元Cable的布置,设计新的三段式布置法,将预应力锚杆(索)在FLAC3D软件层面分为预紧段、自由段、锚固段分别赋予指定参数进行数值计算,建立3组不同支护方式的数值模拟方案对预应力锚固支护效果进行验证。结果表明:预应力锚固后,支护区域顶板受拉应力基本小于破碎岩体顶板抗拉强度,塑性区体积对比普通无预应力支护方案,下降比率达到61.95%,顶板Z方向垂直沉降量降低比率高达83.90%,预应力锚固范围内围岩得到了有效控制。