Gas extraction drilling is a necessary measure for managing gas hazards. For soft coal seams where gas extraction drilling holes are prone to collapse, it is believed that drill rod disturbance is the main cause of hole collapse. This study proposes a research approach to reduce wall stress by optimizing the drill rod structure. Through theoretical analysis, numerical simulation, and industrial tests, a stress model for the drill rod inside the hole was established, and a wall stress equation was derived. The effects of various parameters on wall stress were analyzed. The study suggests optimizing the drill rod structure to reduce the disturbance‐induced wall stress. SolidWorks was used for drilling stress simulation, and a four‐winged concave groove drill rod was developed. After strength verification, comparative industrial tests were conducted. The research results show that as the line density increases, the wall stress of the drilling hole increases. As the length of the suspended section increases, the wall stress initially decreases and then increases. With increasing drilling thrust, wall stress increases linearly, and the growth rate is greater with a larger diameter difference between the drill hole and the drill rod. Numerical simulation results indicate that the critical point maximum stress at the hole entrance, the critical point maximum stress at the hole bottom, and the average stress at the bottom section of the four‐winged concave groove drill rod with a concavity of 5 are significantly reduced compared to those of circular and grooved drill rods. Industrial test results show that using the four‐winged concave groove drill rod significantly reduces the extent of hole collapse. This study provides a reference for addressing the issue of hole collapse in gas extraction drilling for soft coal seams.

In order to solve the problem of drilling deflection, the method of cutting drilling by layers is proposed. The mathematical model of the force of the layered cutting bit was established, and the influencing factors of bit deflection were obtained. The stress equation of the cutting bit is constructed, and the ABAQUS numerical model is established. The influencing factors of the deflection of the bit are analyzed. A comparative industrial test was carried out in the Tunlan Mine. The analysis results show that the main influencing factors of bit deflection are the tangential force on bit f _w , the axial force on bit f _t , the length of the straightener rod l, the elastic modulus of the surrounding rock E, the length of the straightener bit and reaming bit h, and the diameter of the drilling hole D. When the axial force on the bit is between 120 and 220 KN, the change of axial force has almost no effect on the deviation angle of the drill hole. The change of f _w and l is linearly negatively correlated with the bit deviation angle between 50 and 300 mm. Compared with the conventional drill, the horizontal deviation and vertical deviation of the layered cutting bit are reduced by 23% and 43% on average, respectively.

There are a large number of argillaceous weakly cemented strata distributed in western China, and their significant hydraulic characteristics seriously restrict the safe and efficient exploitation of coal resources. In order to solve the problem of deformation and failure of the surrounding rock roadway in this type of strata, this study systematically studied the hydraulic disintegration behavior of argillaceous weakly cemented rocks by comprehensively using laboratory tests, theoretical analysis, numerical simulation and field experiments, established a constitutive model of argillaceous weakly cemented rocks considering hydraulic effects, revealed the deformation and failure mechanism of this type of surrounding rock roadway, and developed the combined support technology of high-strength anchor rod-anchor cable. The technology was applied and verified in the No. 5 mine of Dananhu Lake. The results are as follows: (1) The dried argillaceous weakly cemented rock specimens are more prone to disintegration than those without drying treatment; (2) A mathematical model that can describe the mechanical behavior of argillaceous weakly cemented rocks under hydraulic action is constructed, and the effectiveness of the theoretical model is verified by triaxial compression test. (3) The stress distribution and deformation law of the surrounding rock of the roadway under different moisture content conditions were analyzed: with the increase of moisture content, the stress concentration area of the surrounding rock gradually shifted to the depth, and the extreme value of the principal stress difference increased significantly, resulting in the risk of deformation and instability of the roadway intensified, and the proximity of the top and bottom plate in the middle of the roadway also continued to increase. (4) The joint support parameters are determined by numerical simulation optimization: the diameter of the roof plate and the two anchors is 2400 mm, the length is 2400 mm, and the spacing is 800 × 800 mm; The anchor cable has a diameter of 18.9 mm, a length of 6300 mm, and a spacing of 2400 × 2400 mm. (5) The field test shows that the subsidence of the roadway roof, the amount of two gangs moving closer, the amount of bottom drum and the degree of roof damage are reduced by 35%, 26.1%, 50% and 35.7%, respectively. This study provides a theoretical basis and technical support for the stability control of argillaceous weakly cemented rock roadway.

In western China, there are extensive silty weakly cemented strata, whose significant hydraulic effects severely restrict the safe and efficient mining of coal. In order to address the problem of deformation and failure in roadways within weakly cemented muddy surrounding rock, This paper investigates through indoor experiments, theoretical analysis, numerical simulation, and field tests, The study investigated the hydro-mechanical disintegration of clayey weakly cemented rocks, established the constitutive relationship of clayey weakly cemented rocks under hydraulic influence, determined the deformation and failure process of surrounding rock in tunnels of clayey weakly cemented formations, proposed a high-strength bolt (cable) combined support technology, and conducted field tests at the No. 5 Mine of DaNanHu. The research results indicate:①Under soaking conditions, the undried muddy weakly cemented rock blocks begin to absorb water and crumble after 1 hour, and disintegrate into smaller pieces after 1.5 hours. After drying, the clayey weakly cemented rock samples will exhibit muddying and disintegration phenomena after 20 minutes. After 3 hours, the samples disintegrate into larger blocks, completely losing their strength. ②A mathematical model of muddy weakly cemented rock under hydraulic action was established, and its validity was demonstrated through comparison with triaxial compression tests. ③When the rock water content increases from 2% to 6%, the distance from the stress concentration zone in the surrounding rock of the roadway to the roadway surface increases from 3.1 m to 6.7 m. When the moisture content is in the range of 8–12%, the extreme value of the principal stress difference increases significantly with the increase of moisture content, aggravating the deformation and instability risk of the roadway. When the moisture content reaches 12%, the peak values of the principal stress difference in the roof and ribs are approximately 13.0 MPa and 11.8 MPa, respectively. When the moisture content increases from 2% to 12%, the convergence of the roof and floor in the middle of the roadway increases from 410mm to 2110mm, and the convergence of the two sidewalls in the middle of the roadway increases from 374mm to 1020mm. ④The joint support parameters were determined through simulation tests: The roof and side walls use anchor bolts with a diameter of 22mm and a length of 2400mm, with a spacing of 800×800mm between the bolts. The anchor cables have a diameter of 18.9mm and a length of 6300mm, with a spacing of 2400×2400mm between the cables. ⑤The support parameters determined based on simulation tests were field-tested in the return air tunnel of the 1801 working face at DaNanHu No. 5 Mine.Using high-strength anchor (cable) combined support technology, the maximum settlement of the roadway roof is 140mm; the maximum separation measured between base points located 3.5m, 4.5m, and 5.5m from the roadway roof is approximately 25mm; The maximum separation measured between base points at 0.5 m, 1.5 m, and 2.5 m from the tunnel roof was about 30 mm. The axial stable loads of roof bolts 1 and 2 and side bolts were 90 kN, 120 kN, and 60 kN, respectively, all below the failure load of the bolts, effectively controlling the deformation and damage of the mudstone weakly cemented tunnel.