下向钻孔机械破煤造穴快速卸压增透机制及瓦斯抽采技术研究.pdf

博士学位论文 下向钻孔机械破煤造穴快速卸压增透机制 及瓦斯抽采技术研究 Research on the Mechanism of Quick Stress Relief for Enhancing Permeability by Mechanical Cavitation in Downward Boreholes and Gas Drainage Technology 作者郝从猛 导师程远平 教授 中国矿业大学 二〇二一年五月 万方数据 学位论文使用授权声明学位论文使用授权声明 本人完全了解中国矿业大学有关保留、使用学位论文的规定，同意本人所撰 写的学位论文的使用授权按照学校的管理规定处理 作为申请学位的条件之一， 学位论文著作权拥有者须授权所在学校拥有学位 论文的部分使用权，即①学校档案馆和图书馆有权保留学位论文的纸质版和电 子版，可以使用影印、缩印或扫描等复制手段保存和汇编学位论文；②为教学和 科研目的，学校档案馆和图书馆可以将公开的学位论文作为资料在档案馆、图书 馆等场所或在校园网上供校内师生阅读、浏览。另外，根据有关法规，同意中国 国家图书馆保存研究生学位论文。 （保密的学位论文在解密后适用本授权书） 。 作者签名 导师签名 年 月 日 年 月 日 万方数据 中图分类号 TD7 学校代码 10290 UDC 622 密 级 公开 中国矿业大学 博士学位论文 下向钻孔机械破煤造穴快速卸压增透机制及 瓦斯抽采技术研究 Research on the Mechanism of Quick Stress Relief for Enhancing Permeability by Mechanical Cavitation in Downward Boreholes and Gas Drainage Technology 作 者 郝从猛 导 师 程远平 申请学位 工学博士学位 培养单位 安全工程学院 学科专业 安全科学与工程 研究方向 煤矿瓦斯防治 答辩委员会主席 评 阅 人 二○二一年五月 万方数据 致谢致谢 时光荏苒，犹如白驹过隙。转眼间六年的矿大生活即将结束，回首过往，不 免思绪万千，留恋之情亦油然而生。抚今追昔，愈加清晰的是二十余年的漫漫求 学路即将走到尽头，而我也不再是村庄里的那名懵懂少年。一路走来，有过成功 的喜悦，有过失败的沮丧，更有太多的人和事值得记忆和感动。谨借此论文完成 之际，向给予我帮助和鼓励的恩师益友致以最诚挚的谢意。 衷心感谢我的恩师程远平教授对我的谆谆教诲和悉心指导，本论文从选题、 试验方案设计及实施、现场验证，到论文的撰写与修改，各个环节无不凝聚着程 老师的心血与汗水。读博期间，程老师敏锐的洞察力、一丝不苟的治学作风、严 谨的治学态度、 忘我的工作精神， 以及对科研事业孜孜不倦的追求深深感染着我， 使我懂得作为一名科研工作者应该拥有怎样的精神，应该守住什么样的初心师 恩如海， 衔草难报。 唯有在今后的工作中谨遵恩师教诲， 做一个有品德、 有技术、 有良知的科技工作者，为科研事业的发展贡献自己的绵薄之力。 感谢课题组王海峰、王亮、周红星、刘洪永、李伟、吴冬梅、蒋静宇、刘清 泉等老师在科研及生活上给予的指导和帮助 感谢同窗挚友王振洋博士和商政博 士，同甘共苦一起成长的经历是我们珍贵的回忆感谢田富超师兄给予我的鼓励 和帮助感谢课题组金侃、陈明义、郭海军、赵伟、董骏、涂庆毅、张荣、张浩、 朱金佗、蔺甲、程铭、张开仲、刘正东、戚雨霄、陈二涛、张强、苏二磊、聂雷、 胡彪、蒋炤南、张兴华、王德洋、雷洋、王成浩、易明浩、陆壮、赵长鑫、郝志 远、赵辉、李龙飞等师兄弟在科研、试验、资料收集等方面给予的无私帮助 感谢河南铁福来装备制造集团股份有限公司董事长赵玉凤女士在项目开展 过程中给予的大力支持，在为人处事和生活上给予的指导、帮助感谢公司总工 程师武国胜、总经理武泽铭、技术总监徐云辉、销售总监李静波、研发部长高彬 彬、部长刘路伸、胡金星、王进保等在试验方面给予的支持和帮助。 感谢平煤集团张建国总工程师、张晋京处长、李喜员所长、郭明功总工等领 导对现场试验给予的指导和支持 感谢黄春明处长对我人生观方面带来的启发及 生活上的照顾 感谢父母、岳父岳母和我的妻子张彦飞女士，正是他们的理解、支持、关怀 和包容给了我勇往直前的勇气和信心 感谢哥哥、 姐姐、 弟弟为家庭的辛苦付出， 让我无后顾之忧，安心读书，顺利完成学业，手足情深，没齿难忘。 感谢论文中所引文献的各位作者 限于水平有限，论文中难免有不足之处，恳请各位专家给予指正。最后，衷 心感谢各位专家在百忙中评阅本论文。 万方数据 I 摘摘 要要 顶板巷瓦斯抽采作为突出煤层瓦斯治理的重要方法， 不仅可以通过施工下向 钻孔进行条带瓦斯治理，而且还是工作面回采期间采空区瓦斯治理的有效措施， 具有“一巷两用”的作用。然而，由于缺少便捷高效的卸压措施，顶板巷中主要 通过施工下向密集钻孔进行瓦斯治理。 为解决顶板巷中难以开展高效卸压增透措 施的难题， 本文以平顶山矿区为研究对象， 基于对现场数据和实验室试验的分析， 结合理论研究得到了高应力低渗煤体瓦斯高效抽采途径和卸荷行为对煤体损伤 破坏及增透影响的力学机制； 根据下向钻孔破煤造穴技术困境，论证了新型机械 造穴技术在淹没环境下的破煤优势、破煤过程及受力特征，并基于理论分析获得 了下向钻孔输煤排渣特征；根据机械造穴相似模拟实验和数值模拟分析， 获得了 下向钻孔机械造穴刀具的破煤效果、造穴煤体的卸荷损伤及增透特征；最后，根 据现场试验建立了下向钻孔机械造穴技术体系， 并通过系统的效果考察获得了下 向钻孔机械造穴煤体强化瓦斯抽采效果。本文的主要结论如下 （1）平顶山矿区东西部矿井的瓦斯地质情况差别较大，东部矿井最大主应 力为49 MPa，最大瓦斯压力为3.5 MPa，最大瓦斯含量为27 m3/t，比西部矿井地 应力约高27 MPa，瓦斯压力约高0.82.0 MPa，瓦斯含量约高510 m3/t，而同一区 域内相同埋深条件下，己组煤的瓦斯压力和瓦斯含量比戊组煤分别约高0.7 MPa 和6 m3/t，突出危险性呈现东部高于西部、己组煤高于戊组煤的特点；结合典型 突出矿井的工作面瓦斯治理模式发现， 在瓦斯压力和瓦斯含量相对较低的戊组煤 和西部矿井的己组煤中多采用顶板巷治理瓦斯， 而东部矿井己组煤中多采用底板 巷治理瓦斯，表明顶板巷在以卸应力为主兼顾抽采瓦斯的煤层中具有一定的优 势。 同一煤层不同埋深煤样的多元物性参数测定结果表明，两组煤样的煤质特征 及孔裂隙结构差异不明显，因此，应力环境不同是导致其瓦斯抽采效率差异的主 要原因，在此基础上建立了考虑应力响应的渗透率演化模型，并结合实测渗透率 随埋深变化情况论证了卸荷是实现高应力低渗煤层高效瓦斯抽采的根本途径。 （2）初始围压分别为5 MPa、10 MPa和15 MPa时，卸围压（25 N/s）加轴压 路径下煤样的峰值应力分别是定围压加轴压时的41.4、29.0和34.3，对应的 煤样破坏后的渗透率突增倍数从119.1倍、 75.2倍和86.8倍提高到了308.4倍、 272.6 倍和183倍，表明卸围压条件下煤体更容易破坏并产生更加显著的增透效果；而 以50 N/s卸围压加轴压条件的煤样峰值应力分别是以25 N/s卸围压加轴压时的 77.7、77.6和62.2，煤样破坏后的渗透率增加倍数从308.4倍、272.6倍和183 倍提高到了340.6倍、314.9倍和342.9倍，说明损伤对提高渗透率具有直接显著的 效果， 而且增透效果随着卸荷速率的增大而增大。 另外， 静水压30 MPa降到2 MPa 万方数据 II 过程中煤体渗透率提高了51倍，说明只卸荷也能够有效提高煤体渗透率，但效果 明显低于卸荷后损伤的煤体。 （3）对传统水力造穴技术和新型机械造穴技术在下向钻孔环境下的破煤深 度和破煤体积的分析结果表明在淹没环境下水射流传播速度显著降低，随着水 射流速度的增加虽然破煤深度有所增加，但效果并不显著，而机械造穴的破煤过 程不受淹没环境影响。在相同时间下，机械造穴刀具的破煤深度比不同速度的水 力破煤（170 m/s、190 m/s和210 m/s）提高了5.8倍、4.9倍和4.2倍；在相同的推进 距离条件下， 机械造穴刀具的破煤体积比不同速度的水力破煤 （170 m/s、 190 m/s 和210 m/s）提高了9.7倍、7.8倍和6.3倍，两种造穴技术的破煤效率差异充分证明 了机械破煤造穴技术明显优于水射流破煤。 （4）机械造穴相似模拟实验表明，机械造穴刀具张开过程分为两个阶段， 第一个阶段和第二阶段分别以6.1和46.3的扩张角扩大，并在第二阶段快速张开 将孔径扩大到500 mm，同时，根据钻机扭矩调整实验认为造穴过程中的推进速 度以不超过钻进速度的20为宜。 结合相似实验结果开展了造穴煤体损伤增透数 值模拟分析，结果表明造穴后煤体径向应力卸压范围从1.3 m增加到6.2 m，提 高了4.8倍；最大塑性破坏范围从0.3 m增加到3.75 m，提高了12.5倍；钻孔周围煤 体渗透率提高10倍的范围从0.95 m增大到6 m，提高了6.3倍；抽采30180 d的有 效半径提高了1.942.14倍。 （5）根据现场试验确定了下向钻孔机械造穴过程的施工参数（推进压力8 MPa、 旋转速度90 r/min、 推进速度0.2 m/s） 和排渣参数 （泵站流量550600 L/min） ； 在此基础上开展了系统的现场应用和效果考察，结果表明，机械造穴段钻孔出煤 量约为262 kg/m，大于理论出煤量255 kg/m，说明机械造穴较好的达到了设计直 径500 mm；煤层渗透率从造穴前的0.0018 mD提高到造穴后的0.0431 mD，增加 了23.9倍；初始钻孔百米瓦斯纯量从造穴前的0.36 m3/min hm提高到造穴后的 2.1 m3/min hm，提高了5.8倍；在造穴钻孔比普通钻孔数量减少70的前提下， 瓦斯抽采达标预抽期从90 d降低到70 d；造穴钻孔预抽瓦斯结束后，巷道掘进速 度从4.2 m/d提高到4.6 m/d，最大钻屑量从4.5 kg/m降低到3.9 kg/m，掘进期间各 项指标均明显低于临界值。 该论文有图126幅，表27个，参考文献184篇。 关键词关键词顶板巷；机械造穴；卸压增透；瓦斯抽采；水力冲孔 万方数据 III Abstract Roadways gas extraction that are above the roof play key roles in outburst prevention process. On one hand gas pre-drainage boreholes can be drilled from there and on the other hand, the boreholes can be used for goaf gas drainage. These two benefits are obvious. However, due to the lack of convenient and efficient pressure relief measures, gas control in roof roadway is mainly carried out through construction of dense boreholes. In order to solve the problems that are encountered during the gas drainage process, in this paper, we are based on the Pingdingshan coal area. Through field observation, laboratory analysis and model development, the high efficiency gas drainage means and coal stress relief mechanism are obtained. According to the technical difficulties of downhole coal cavitation, the advantages of coal breaking, coal breaking process and force characteristics of the new mechanical cavitation technology in submerged environment were demonstrated. Based on the similarity simulation experiment and numerical simulation analysis of mechanical cavitation, the coal breaking characteristics of the downhole mechanical cavitation tool and the damage and anti-reflection characteristics of the cavitation coal are obtained. At last, a system is developed based on the field trials of downward borehole coal breakage and it is verified through field production. The main findings are as follows 1 Statistical analysis found that the gas geology of the east and west mines in Pingdingshan mining area is quite different. In the eastern mine, the maximum principal stress is 49 MPa, the maximum gas pressure is 3.5 MPa, and the maximum gas content is 27 m3/t. Compared with the western mine, the ground stress is about 27 MPa higher, the gas pressure is about 0.82.0 MPa higher, and the gas content is about 510 m3/t higher. Under the same buried depth in the same area, the gas pressure and gas content of coal of group Ji are about 0.7 MPa and 6 m3 /t higher than those of group Wu, respectively. The outburst risk is higher in the east than in the west and higher in Ji group of coal than that in Wu. Combining with the gas control mode of typical outburst mines, it is found that roof roadway is used for gas control in coal of Group Wu and Group Ji in western mine, where gas pressure and gas content are relatively low. In the eastern coal mines, bottom-drawing roadways are often used to control gas. That shows that the roof roadway has certain advantages in coal seams that focus on stress relief while taking into account gas extraction. The results show that the difference of coal quality and pore fracture structure between two groups of coal samples with different 万方数据 IV buried depth is not obvious. Therefore, the different stress environment is the main reason for the difference in gas drainage efficiency. On this basis, a permeability evolution model considering the stress response is established, and combining with the measured permeability variation with the buried depth, it is proved that unloading is the fundamental way to realize efficient gas extraction in high-stress and low-permeability coal seams. 2 When the initial confining pressure is 5 MPa, 10 MPa, and 15 MPa, the peak strength of coal going through relieving confining pressure 25 N/s and adding axial pressure is 41.4, 29.0, and 34.3 of that of the force path of constant confining pressure and adding axial pressure. And correspondingly, the permeability increases from 119.1 times, 75.2 times and 86.8 times to 308.4 times, 272.6 times and 183 times after coal sample destruction. The results show that under the condition of discharging confining pressure, the coal is more prone to damage and produces higher anti- reflection effect. However, the peak strength of coal going through relieving confining pressure 50 N/s and adding axial pressure is 77.7, 77.6 and 62.2 of that under the condition of confining pressure 25 N/s and adding axial pressure. And the permeability increases from 308.4 times, 272.6 times and 183 times to 340.6 times, 314.9 times and 342.9 times after coal sample destruction. That means the faster the unloading rate, the more significant the permeability increase. The results of coal permeability measurement with helium show that when the hydrostatic pressure drops from 30 MPa to 2 MPa, the coal permeability increases by 51 times, indicating that unloading can effectively increase the coal permeability. 3 The analysis results of coal breaking depth and coal breaking volume under the downhole environment with traditional hydraulic cavitation technology and new mechanical cavitation technology show that The efflux propagation velocity decreases significantly in submerged environment, and with the increase of water jet velocity, the coal breaking depth increases, but the effect is not significant. However, the process of mechanical coal breaking is not affected by the submerged environment. At the same time, the coal breaking depth by mechanical cavitation tool is 5.8 times, 4.9 times and 4.2 times higher than that of hydraulic cavitation at different speeds 170 m/s, 190 m/s and 210 m/s. Under the same advancing distance condition, the coal volume of mechanical cavitation tool is 9.7, 7.8 and 6.3 times higher than that of hydraulic at different speeds 170 m/s, 190 m/s and 210 m/s. The difference of efficiency between the two technologies fully proves that mechanical coal breaking cavitation technology 万方数据 V is obviously superior to water jet coal breaking. 4 Based on the structure of the mechanical cavitation cutter, the stress characteristics of the cutter during coal breaking process were obtained by numerical simulation. Combined with similarity simulation experiment of mechanical cavity technology, the splaying process of cavitation cutter can be divided into two stages. The cavitation cutter were splayed with an expansion angle of 6.1 and 46.3 in the first stage and the second stage, respectively, then were fully splayed rapidly in the second stage. Additionally, it is considered that the advance speed during mechanical cavitating should not exceed 20 of the normal drilling speed. The numerical model of coal unloading damage is established combined with similar simulation results. The numerical simulation results shown that the radius of radial stress relief zone increased by 4.8 times from 1.3 m to 6.2 m. The maximum radius of plastic failure zone increased by 12.5 times from 0.3 m to 3.75 m. The range with a ten-fold increase in permeability increased by 6.3 times from 0.95 m to 6 m, and the effective extraction radius within 30180 d increased by 1.942.14 times. 5 The parameters of mechanical cavitation technology propulsion pressure is 8 MPa, rotation speed is 90 r/min, propulsion speed is 0.2 m/s and slag removal pumping flow range 550600 L/min were determined through field tests. In the field investigation, coal weight of the mechanical cavitating borehole section was about 262 kg/m, which is greater than the value of calculation of 255 kg/m, and the diameter of cavity is larger than design value of 500 mm. The coal permeability increased by 23.9 times from 0.0018 mD to 0.0431 mD after cavitation after mechanical cavitating and gas emission per hundred meters of incipient borehole increased by 5.8 times from 0.36 m3/minhm to 2.1 m3/minhm. In addition, the number of mechanical cavitation boreholes reduced by 70 compared with ordinary boreholes and the gas drainage time of reaching expected level was decreased from 90 d to 70 d. After gas drainage, the tunnel excavation rate was increased from 4.2 m/d to 4.6 m/d and the maximum drilling cutting weight was reduced from 4.5 kg/m to 3.9 kg/m, indicating that the risk of coal and gas outburst was significantly reduced during tunneling. This work contains 126 figures, 27 tables, and 184 references. Keywords roof roadway; mechanical cavitation; stress relief and permeability enhancement; gas drainage; hydraulic cavitation 万方数据 VI 目目 录录 摘要摘要 ............................................................................................................................... I 目录目录 ............................................................................................................................ VI 图清单图清单 ........................................................................................................................ XI 表清单表清单 ..................................................................................................................... XIX 变量注释表变量注释表 ............................................................................................................. XXI 1 绪论绪论 ........................................................................................................................... 1 1.1 研究背景与意义..................................................................................................... 1 1.2 国内外研究现状..................................................................................................... 3 1.3 存在的问题............................................................................................................. 7 1.4 主要研究内容和技术路线..................................................................................... 9 2 高应力煤体瓦斯赋存及其流动通道应力响应特征高应力煤体瓦斯赋存及其流动通道应力响应特征 ............................................. 12 2.1 平顶山矿区瓦斯地质特征................................................................................... 12 2.2 煤体多元物性参数及孔裂隙结构特征............................................................... 22 2.3 煤体瓦斯吸附解吸特性....................................................................................... 30 2.4 煤体瓦斯流动通道应力响应特征....................................................................... 33 2.5 深部高应力煤体瓦斯抽采瓶颈及工作面合理增透技术................................... 37 2.6 小结....................................................................................................................... 40 3 卸荷速率对煤体损伤破坏影响的力学机制卸荷速率对煤体损伤破坏影响的力学机制 ......................................................... 42 3.1 实验方法............................................................................................................... 42 3.2 煤样常规压缩实验..................................................................................