Hydraulic permeability increase technology is currently one of the key techniques for gas prevention and control. However, existing methods still suffer from issues such as discontinuous decompression zones and complex construction processes. To further improve coal seam decompression and permeability increase, a novel method of axial plane slotting using in-seam jet cutting is proposed. First, a theoretical model for plane slotting was established, followed by an analysis of how surrounding rock pressure and jet slotting parameters influence the development morphology of pressure-relief fractures. This revealed the pressure-relief and permeability increase mechanism of the axial plane slotting technique along the coal seam, and optimized the axial plane slotting parameters. Field application tests were conducted for three technologies: pre-extraction through boreholes along the coal seam, annular slotting through boreholes to increase permeability, and axial plane slotting through jets to increase permeability. The advantages and disadvantages of the three technologies were compared. The research findings indicate: When the lateral pressure coefficient is 1 and the confining pressure increases from 18.75 MPa to 40 MPa, the fracture development height exhibits a linear upward trend. Below 18.75 MPa, non-through fracture zones exist, and the larger the lateral pressure coefficient approaches 1, the larger the de-pressurized fracture zone becomes. When the horizontal slot spacing is 1.25 m or less, the de-pressurized fracture zone achieves full connectivity. With vertical slot spacing below 5 m, fractures between two rows of slots can form a connected zone. However, when vertical spacing exceeds 5 m, fracture connectivity between slots becomes difficult. For coal seams thicker than approximately 4 m, two rows of slots can be constructed. Field tests indicate that compared to the following-strike drilling pre-extraction technique and longitudinal annular slotting for permeability increase, jet axial planar slotting increases gas extraction concentration by 0.27 and 0.5 times, boosts net gas extraction by 1.43 and 2.35 times, and reduces the time to achieve extraction targets to 48 days. This further validates the effectiveness of the jet axial planar slotting pressure relief and permeability increase technique.

To explore the damage evolution law of gas-bearing coal under impact, a split Hopkinson pressure bar (SHPB) test system for gas-bearing coal was used to conduct dynamic compression tests on coal with gas pressures of 0, 0.5, 1.0, 1.5, and 2.0 MPa. Based on the energy theory, the deformation and failure processes of gas-bearing coal under impact were analyzed, and the influence of gas pressure on energy parameters of coal was discussed. Using the SMP strength criterion and Weibull distribution function, a dynamic damage constitutive model of gas-bearing coal considering gas-impact coupling damage was established by combining the energy consumption index. The results indicate that during the impact compression process, the energy curve of gas-bearing coal can be divided into a slow growth stage, an accelerated growth stage, and a stable stage. With the increase of gas pressure, the reflected energy of coal shows a linear increase trend, while the transmitted energy and dissipated energy show a linear decrease trend. The theoretical curve based on the gas-impact coupling damage constitutive model is highly consistent with the test curve, indicating that the model can accurately describe the damage evolution law of the entire stress-strain process of gas-bearing coal under impact.

The difficulty of underground gas extraction in coal mines depends on the permeability of the coal body, and coal body permeability enhancement is one of the keys to gas extraction research. To improve the permeability of low-permeability coal seams, artificial measures to increase the cracking of coal seams are an effective means to improve the permeability of coal seams. By analyzing the principle of high-pressure water jet seam making, experimental research on the influencing factors of jet flat drag seam making was carried out with a true triaxial test device, and the influencing laws of jet angle, number of drilling holes, jet pressure, ground stress, and seam strength on the seam making effect were analyzed, and on-site applications were carried out. The test results show that the high-pressure water jet can form evenly distributed ‘tooth-like’ cracks on both sides of the slit hole, and with the increase of the jet angle and the increase of the strength of the coal seam, as well as the increase of the ground stress, the length of the cracks is gradually reduced; On the contrary, the crack length becomes progressively longer as the jet pressure increases.Field test results show that the optimal cutting pressure is 80 MPa, when the cutting time is 12 min, the corresponding depth of cutting is about 1.75 m, at this time, the reasonable optimal cutting hole spacing is 6 m, the optimal cutting slot spacing is 4 m. The experimental results can provide a reference for similar coal mine gas extraction and hydrometallisation measures for gas management.

In response to the “three highs” problem in the mining of deep high-gas mines, the rapid increase in the coal seam permeability coefficient and gradual increase in coal and gas outburst problems have made gas control more difficult. This study considered the occurrence of remote outburst coal seams in the Zhujixi Mine as the research background and performed theoretical analysis, calculations, numerical simulations, and other technical methods to analyze the gas occurrence characteristics of the 11-2 coal seam and the feasibility of using this seam as a lower protective layer for mining. The pressure relief protection range for the overlying 13-1 coal seam, to the recovery of the 11-2 coal seam, was determined. A regional anti-outburst technology was proposed for underground through-layer and parallel-layer drilling, focusing on pre-gas extraction for the protective layer. In addition, a pre-gas extraction regional anti-outburst technology combining the surface and underground mining of the protected layer is also proposed. Gas occurrence in the 11-2 coal seam is uneven and has poor regularity, presenting high gas areas. It is significantly affected by the geological structures and shale properties of the coal seam roof and floor. The 11-2 coal seam is a stress-dominated and gas-outburst coal seam. The Zhujixi Mine presents a joint underground extraction and regional outburst prevention mode; that is, the 11-2 coal seam with a lower outburst risk is selected as the protective layer for mining first, whereas the 13-1 coal seam is protected while the gas in the protected layer is extracted. The 11-2 coal is characterized by the gas control mode of “one side, three lanes+ground drillings” to achieve multi-purpose, joint treatment, and continuous mining of one lane. The excavation face exhibits comprehensive anti-outburst measures, such as through-layer drilling pre-extraction and a coal mining face over the layer drilling pre-extraction area. During the mining period, surface drilling and a top extraction roadway are used to extract 13-1 coal-depressurized gas. By adopting joint extraction technology in the upper and lower mining areas, the residual gas content and pressure were measured at the underground excavation and mining working face. The predicted indicators did not exceed the standard levels, and no dynamic phenomena occurred. As a result of the application of the anti-outburst technology in the joint extraction area of the Zhujixi Mine, the proportion of extraction in the upper and lower mining areas was 56.7%, and the proportion of extraction in the underground mining area was 43.3%. These factors are interdependent and indispensable. The maximum height of the caving zone after mining the 11-2 coal face was 11.6 m, whereas the height of the fracture zone was 34.4-52.2 m. The 13-1 protective-layer working face is arranged on the upper part of the fracture zone or lower part of the curved subsidence zone, which can effectively increase the permeability of the 13-1 coal seam. Engineering practice has shown that the joint regional anti-outburst technology and engineering application in Zhujixi mine have achieved good results, forming a regional anti-outburst technology system for joint extraction of mines and providing a reference for the safety production of similar conditions in outburst mines.

To study the stability of roof areas in the process of rapid driving in coal roadways, the deformation law of roof areas under the interaction of different factors is systematically analyzed through mechanical analysis, numerical calculation, and industrial testing, yielding a reasonable selection basis of the critical parameters of roof space stability. The roof is difficult to control when the unsupported roof distance exceeds 2.0 m and the roof thickness is less than 0.9 m. Considering the tunnel excavation safety and effectiveness, the support technology principles of "large-angle control span" and "classification control technology of roadway surrounding rock stability" based on the support "small-step" excavation method were realized in the case study of the 150,802 machine roadway of Liuzhuang Coal Mine, China. The resulting rapid tunneling system made it possible to increase the tunneling speed from 300 to 500 m/month, i.e. by 67%. The whole roadway remained relatively stable, verifying the proposed approach feasibility.

Accurate analysis of failure mechanisms and scientific support for design of coal roadways in water-rich areas plays an important role in the long-term stability control of roadways. In this study, taking the water-rich 1044 return laneway in the Taoyuan Coal Mine of China as an example, laboratory experiments and field tests were performed to study the potential failure mechanisms of the roadway. It was found that the microcracks inside the coal body increase and strength of the coal decreases with the long-term influence of groundwater. The weak surrounding rock and high vertical in-situ stress were the main internal causes of roadway instabilities. Based on the potential failure mechanisms, a new optimal support design including bolt, cable, metal mesh, shotcrete, and grouting was proposed and detailed support parameters were introduced. A field experiment was performed with the new support, and surface deformations of the experimental roadway were monitored. Results show that the new support can reduce the deformation of the 1044 return laneway by nearly 90% compared with those of the original design. Support technology presented in this contribution provides a significant reference for the control of roadways in water-rich areas.

针对松软煤层顺煤层定向瓦斯治理钻孔打不深、打不准、成孔率低的难题,详细分析了现有顺层定向钻进的技术瓶颈,创新尝试使用了底板定向梳状条带消突钻孔,突破了主孔快速开分支、分支孔水力下筛管护孔等技术难题,并选取杨柳煤矿1077中段机巷进行现场试验,在煤层坚固性系数为0.46的松软层中,共完成总进尺8 494 m,实验效果达到设计要求.现场试验表明,底板定向梳状条带消突钻孔有效缓解了采掘接替紧张局面,提高了矿井钻孔施工及瓦斯治理技术水平,为松软煤层定向瓦斯治理钻孔施工提供了新的技术思路.

Coal and gas outburst is an extremely complex dynamic phenomenon of mine gas, which is mainly manifested in a very short time. A large amount of coal and rock are thrown out from the coal body to the mining space and accompanied by a large volume of high-pressure gas. In the process of coal and gas outburst, the internal energy consumption of gas is composed of two parts: one is used to throw out broken coal and rock mass, and the other is used to pulverize broken coal. In this article, from the perspective of energy dissipation, the experiment of broken coal ejection with different coal particle sizes, different adsorption gas, and pressure is studied. The characteristics of coal ejection are studied and analyzed, and the proportion of adsorbed gas participating in the outburst work is quantitatively analyzed. The results show that after outburst excitation, residual gas will continue to desorb and work on outburst until the power is insufficient to throw coal body; compared with air, CO2 gas has a stronger ability to work on the outburst, and the outburst coal is thrown far away, and the pulverization effect is stronger. Through the energy analysis in the process of outburst, the results show that when the particle size of the coal sample is consistent, the greater the outburst pressure is, the larger the desorption amount of the adsorbed gas is, and the larger the volume involved in the outburst work is. When the test gas is consistent with the outburst pressure, the gas desorption amount of the small-size coal sample is more, the desorption gas has a stronger ability to work on the outburst, and the proportion of participating in outburst work is higher. The crushing degree of coal plays an important role in the expansion and release of gas internal energy.

Aiming at the problem of the deformation of the roadway floor plate during the laneway during the retention period, the mechanical model of the roadway floor is established, and the deformation characteristics of the roadway floor and the change law of the bottom drum are studied and analyzed through theoretical calculation and calculus simulation, revealing the instability mechanism of the surrounding rock of the roadway under the stress disturbance environment, and when not affected by the adoption, the roadway forms a certain stress concentration area within the effective range of support. During mining, under the comprehensive action of the original peripheral stress field and the mining stress field, the cliffhanger is unstable under the comprehensive action of the original peripheral stress field and the mining stress field, and the extrusion and stretching effect of the unflapped part of the rock layer above the goaf section of the coal seam is set up along the air, resulting in violent deformation such as the bottom drum, and the rotational sinking of this part of the unflinted rock layer further aggravates the transfer of the overburden load to the surrounding rock of the lane, so that the surrounding rock along the empty lane is subjected to a large additional stress, and the mining stress field plays a leading role, and the mining stress “far field” is the compound stress field, of which the tensile stress is the leading destructive factor. The deformation of the surrounding rock is mainly based on the bottom, and the horizontal stress on the bottom plate along the empty lane is mainly generated by the horizontal strain that occurs after the lower rock layer of the filling body and the coal gang is subjected to the supporting pressure transmitted by the top plate. With the mining of the working surface, the roof of the goaf area is broken and collapsed to form the characteristics of “vertical three belts,” which is affected by the “large support” of the coal body of the working surface and the “small support” of the surrounding rock along the empty roadway, and the pressure relief of the cut roof can make the roof plate along the empty lane change from the “long arm beam” structure when the roof is not cut into the “short arm beam” structure, blocking the lateral stress of the goaf area to the roof plate of the alley and significantly reducing the degree of stress superposition of the roof plate of the alley. The technical means of blasting cutting roof active pressure relief and protective lane are used to block the transmission of lateral support pressure, the roof slate layer is precracked in advance, the sinking of the rock layer is accelerated, the disturbance time is reduced, the vertical stress of the rock layer and the rock layer above it along the empty roadway is reduced, the vertical stress concentration of the roadway is reduced, the stress concentration coefficient is reduced, the degree of damage of the surrounding rock after the top is weakened, the damage range is reduced, and the technical problem of large deformation prevention and control along the bottom drum of the empty alley can be solved. Constructing the mechanical structure model of the top plate of the cut top pressure relief and the uncut top pressure relief along the empty lane, the stress change characteristics of the active protective rock surrounding rock along the hollow top of the cut top pressure relief were calculated, and after the technical scheme of the blasting cut top active pressure relief and protection lane was adopted, the deformation along the empty roadway was significantly weakened, the stability of the surrounding rock of the roadway after the blasting of the cut roof was significantly improved, the maintenance state along the section of the empty roadway was good, and the cross-sectional convergence rate was reduced by 37.3% compared with the original section. Cutting the roof active pressure relief and protective lane can effectively improve the stability of the surrounding rock.

为了探索承压水下工作面自动开采模拟实验及覆岩运移应力监测,通过对承压含水层与采煤支护系统模拟设计的结合,设计了一种承压水下自动采煤相似模拟实验系统及方法,实现承压水下的相似模拟模型的自动开挖和压力监测,包含模拟承压含水层和采煤支护模拟系统.设计的模拟承压含水层,通过调节水箱高度和调压阀实现对承压含水层水压力的控制,以及水袋之间的水流量传感器,实现在采动过程中对含水层中受采动影响部分的水的流动及补给规律进行模拟及监控.设计的采煤支护模拟系统,实现在上覆岩层中存在含水层时,对承压水下煤层开采的近似模拟与覆岩运移应力监测,实现了自动开挖模型功能,提高了实验准确性,大大降低工作强度,使实验更加智能化.

Abstract In view of the”three high” problem of deep high gas mine mining, the rapid increase of coal seam permeability coefficient and the continuous reduction of coal and gas outburst risk gradually increase the difficulty of gas control. Taking the occurrence of long-distance outburst coal seams in zhujixi mine as the research background, this paper analyzes the gas occurrence characteristics of 11 − 2 Coal Seam and the feasibility of 11 − 2 Coal Seam as the lower protective layer by using the technical methods of theoretical analysis, calculation and numerical simulation, and studies and determines the pressure relief protection range of 11 − 2 coal seam mining to the overlying 13 − 1 coal seam; This paper puts forward the regional outburst prevention technology of underground penetrating and bedding borehole pre drainage for the protective layer and the regional outburst prevention technology of pre drainage combined with the mining surface and underground of the protected layer. Coal seam 11 − 2 has uneven gas occurrence and poor regularity. There are only high gas areas in some parts, which are greatly affected by geological structure and slate properties of coal seam top and bottom. Coal seam 11 − 2 is a stress dominated briquette and gas outburst coal seam. Zhujixi mine adopts the regional outburst prevention model of combined pumping up and down the well, that is, coal seam 11 − 2 with low outburst risk is selected as the protective layer for mining first, and coal seam 13 − 1 is protected above, At the same time, the gas in the protected layer shall be extracted. 11 − 2 Coal adopts the gas control mode of "one side with three lanes (one side with five lanes in the first mining face) and surface drilling", so as to realize multi-purpose, joint treatment and continuous mining of one lane. Comprehensive outburst prevention measures are adopted for the tunneling working face through layer drilling pre drainage and the coal mining working face along layer drilling pre drainage area. During the mining period, the surface drilling and top drainage roadway are used to extract the pressure relief gas of 13 − 1 coal, and the combined upper and lower shaft extraction area technology is adopted. In the underground excavation and mining working face, the gas concentration of return air flow, measured residual gas content and residual gas pressure are measured, The prediction indexes have not exceeded the standard, and there has been no dynamic phenomenon. In the application of outburst prevention technology in the area of up-down combined extraction in zhujixi mine, the proportion of well extraction is 56.7%, and the proportion of underground extraction is 43.3%. The maximum height of the caving zone after mining of coal face 11 − 2 is 11.6m, and the height of the fracture zone is 34.4 ~ 52.2m. The working face of protective layer 13 − 1 is arranged at the upper part of the fracture zone or the lower part of the bending subsidence zone, which can effectively increase the permeability of coal seam 13 − 1, and has no significant impact on the normal mining of coal face 13 − 1.

Water-based SiO2 nanofluid, a wetting agent, has been proved to be feasible in improving the control effect of coal seam water injection against coal and gas outburst. However, the effect of nanofluid concentration on the control effect is not clear. Therefore, it is necessary to investigate the wettability-enhancing effect of nanofluid with different concentrations. In this paper, firstly, the effect of SiO2 nanofluid with different concentrations on water adsorption capacity of coal was obtained. Then, the wettability-enhancing effect of SiO2 nanofluid with different concentrations was investigated by contact angle tests. Finally, the effect of nanofluid concentration on the wettability enhancement of coal was revealed. The results showed that increasing the nanofluid concentration can enhance the water adsorption capacity and wettability of coal. In addition, the wettability-enhancing effect of nanofluid gradually got worse with time, while the stability of wettability-enhancing effect gradually got better. Furthermore, the 0.5 wt% nanofluid had the best sustainability and stability of the wettability-enhancing effect, while 1.0 wt% had the worst sustainability and stability. The research achievements could provide important basis for the selection of nanofluid concentrations in coal seam water injection.

In order to solve the problems of the uneven deformation of Gangue Filled Wall and the difficulty of large-scale promotion of roadway side support, and to achieve the purposes of direct disposal of coal mine waste, reducing costs, and protecting the environment, the failure mechanics model of the bagged gangue was established, and the mechanical action relationship between longitudinal external load and transverse external load of gangue woven bag was deduced. Through the uniaxial compression test of large-scale flexible backfill (coal gangue of different particle sizes), it was obtained that when the strain is 0.2, the bearing capacity of particles with particle sizes between 0 and 10 mm is greater than 5 MPa, and when the strain is 1.27, the bearing capacity of particles with particle thicknesses between 10 and 20 mm is greater than 0 mpa, which meets the requirements of resistance value and resistance growth rate of gob side entry. In the “load deflection” test of backfill (gangue) samples, it was found that the maximum failure load of wet shotcrete is greater than that of dry shotcrete, and the wet shotcrete can withstand greater deformation under the same load conditions. Through the analysis of the experimental results of “flexural strength thickness” and “maximum failure load thickness”, it was finally determined that the thickness of the spray layer with good flexibility and sufficient support force is controlled at about 80 mm.

瓦斯灾害是煤矿安全生产的"第一杀手",瓦斯抽采是防治瓦斯事故的治本之策,而低透气性煤层是制约瓦斯抽采的关键因素.由于浅部资源逐渐枯竭,我国煤矿开采深度以每年10～30, m速度向深部延深,开采煤层表现出显著的"低煤层渗透性、高围岩地应力、高煤层瓦斯压力"特征.面对深部矿井煤岩复合动力灾害,单一采用地面井或者井下钻孔预抽等措施,在高地应力、低渗透性条件下难以取得较为理想的抽采效果,同时无法消除煤岩内部积聚的弹性应变能、降低采掘区域内地应力威胁,需结合有效的卸压增透措施才能满足复合动力灾害的治理需求.由此,研究一种既能满足煤层瓦斯快速抽采达标,又能达到降低煤层应力的技术手段,是煤矿科研工作者共同努力的方向.

To study the evolution law of axial force and shear stress of a full-length anchorage bolt in a rectangular roadway during roadway driving and working face mining, based on the stress analysis of the bolt, considering the elastic parameters and geometric size of the bolt, the effect of a bearing plate on surrounding rock, roadway cross-section shape, roadway deformation degree, and roadway elastic parameters, elastic mechanics and mathematical analysis methods were used to establish the mechanical model describing the interaction between the bolt and surrounding rock, and the mechanical formulas for calculating the axial force and shear stress of the bolt were derived. Taking the mining roadway of 1,131(1) working face in the Zhujidong coal mine of the Huainan mining area as the engineering background, the axial force and shear stress of the bolt in the middle of the roof and side of the rectangular roadway with the advance of driving face and working face were analyzed. The mechanical model and theoretical analysis results are verified by installing force measuring bolts with the same mechanical properties as the field and observing the real axial force distribution of the bolts.

Ultra-high-pressure hydraulic slotting technology is an effective method to realize pressure relief and permeability enhancement of a single coal seam. In this paper, through a hydraulic slotting borehole-based total gas extraction amount calculation model, it was concluded that ultra-high-pressure hydraulic slotting technology could increase the gas extraction surface area, improve the gas flow mechanism, and transform single-layer radial flow into interlayer, radial composite flow, thus greatly enhancing the gas extraction efficiency. Based on theoretical results, a field test was conducted in the Renlou coal mine, Anhui Province, China. According to the actual characteristics of the 7 2 coal seam in the Renlou coal mine, the key slotting parameters of this coal seam were determined. The gas extraction sources of each coal seam could be determined accurately through multi-gas source identification and tracer technology, and the effect of hydraulic slotting pressure relief and permeability enhancement was investigated. The test results indicated that the 100-day gas extraction concentration in boreholes using ultra-high-pressure hydraulic slotting technology was 2.68–7.59 times that in conventional boreholes. The average daily pure extraction volume of slotting boreholes was 3.94 times that of conventional boreholes. Based on the calculation of the gas extraction radius of boreholes in the slotted area, it was found that, under the condition of meeting the same control range of outburst elimination, drilling work could be reduced by more than two-thirds. These achievements could provide critical references for the application of ultra-high-pressure hydraulic slotting technology.

There is a "bottleneck effect" in the extraction process of ordinary boreholes. It is crucial to enhance the permeability of the boreholes. In this paper, a set of ultra-high-pressure water jet-based drilling and slotting integrated equipment was designed and applied, and the ultra-high-pressure safeguarding technology of several key components has been studied. Finally, the gas drainage effect of slotted boreholes and ordinary boreholes was investigated through field tests. The test results showed that after using ultra-high-pressure hydraulic slotting, the average drainage concentration of slotted boreholes is 1.49 times that of the ordinary boreholes. The average extraction scalar volume of slotted drilling is 3.02 times that of ordinary boreholes. Under the condition of original content of 12m(3)/t, the extraction radii of slotted boreholes in 3 months and 6 months are 3.76m and 4.76 m, respectively. The effective radius of the slotted boreholes is 1.76 times that of the ordinary boreholes, which indicate that the ultra-high-pressure hydraulic slotting technology can effectively relieve the pressure and increase the permeability of the soft coal seam.

In order to reduce the risk of coal and rock dynamic disasters in the coal mine production process, the coupling mechanics characteristics of coal and rock produced in the process of coal mining in the Dingji Coal Mine are taken as the research object, and the experimental study on the deformation characteristics and the variation rule of mechanical parameters of raw coal under multifield coupling (temperature, gas, and stress coupling) was carried out. The results show that the elastic modulus, peak strain, and peak stress of raw coal samples under the thermal-hydraulic-mechanical coupling have the same change law in the test temperature range and all of them show a linear decreasing law as the temperature increases. Under the same temperature gradient increasing condition, the elastic modulus, peak strain, and peak stress show a nongradient decreasing trend as the temperature increases. Both the deformation modulus and the lateral expansion coefficient show a linear increase as the temperature increases, while the deformation modulus and the lateral expansion coefficient show a nongradient increase trend as the temperature increases under the same temperature gradient increasing condition. Under the action of the thermal-hydraulic-mechanical coupling, unloading confining pressure obviously accelerated the yield process of the coal sample, and the confining capacity of confining pressure on transverse strain of the coal sample decreased. To prevent the occurrence of coal and gas outburst, it is necessary to take specific prevention measures according to the change law of triaxial compression mechanics of a raw coal specimen under the action of the thermal-hydraulic-mechanical coupling.

Aiming at the problem of coal and gas outburst prevention and control in serious outburst dangerous coal seam under complex geological conditions in Qidong Coal Mine, the water jet drilling and expanding integrated device was used to implement the measures that expand holes and increase permeability by high-pressure water at a 912 working face. The experimental results show that when the pump pressure is between 13 MPa and 15 MPa, the particle size of pulverized coal can be effectively reduced and the discharge of pulverized coal can be increased. The coal output of a single hole is 0.8~2.3 t/m, with an average of 1.1 m/t. The unreamed conventional borehole has a single-hole initial gas drainage rate of 50.74 m3/d and a single-hole cumulative gas drainage rate of 736.12 m3 within 30 days. The initial single-hole gas drainage volume of the reaming borehole is 178.74 m3/d, and the cumulative single-hole gas drainage volume in 30 days is 2227.57 m3. After reaming, the initial gas drainage scalar of a single hole increases 3.52 times of the original unreamed conventional drilling and the average cumulative gas drainage scalar of a single hole reaches 3.03 times of the original unreamed conventional drilling. This technology and equipment have a good promotion and application value in gas treatment of soft, low-permeable coal seams.

为提高顺层瓦斯抽采钻孔封孔效果,以贵州五轮山煤矿8#煤层地质条件为工程背景,开展合理的封孔深度及注浆参数研究.通过分析钻孔漏气影响因素,将钻孔漏气形式分为巷道裂隙带漏气、钻孔裂隙带漏气、孔壁边缘漏气、封孔段材料漏气4类.采用理论分析、数值模拟及现场试验的方法,对巷道及钻孔周围应力分布规律开展研究,确定合理封孔深度为15 m,合理注浆扩散半径为0.40～0.50 m.研发了新型封孔材料,其抗压强度提升了20 MPa、膨胀率增加了20％、黏度降低了23％、终凝时间缩短了5h.基于注浆柱面扩散理论,确定合理注浆压力为1.5 MPa,合理注浆量为0.3 m3.在1811工作面运输巷道布置16个顺层瓦斯抽采钻孔,开展4组不同封孔参数条件下的工业性试验,结果表明:在封孔深度为15 m、注浆压力为1.5 MPa、注浆量为0.3 m3条件下封孔效果较好、经济性较优,钻孔日平均抽采瓦斯浓度提升30％.