Rock burst mine mechanical reaming unloading-supporting coupling anti-impact method and rock burst prevention and treatment structure

Abstract

The invention discloses a rock burst mine mechanical reaming unloading-supporting coupling anti-impact method, and belongs to the technical field of coal mine rock burst prevention and control. The method comprises the steps of firstly collecting geological data and evaluating impact risk; then calculating parameters and constructing a conventional pressure relief drill hole comprising a shallow hole section and a deep hole section; then mechanical reaming is conducted on the deep hole section, a large-diameter pressure relief space is formed, and elastic coal pillars between holes are reserved; and finally, dense shallow holes are complementarily drilled among the reaming drill holes step by step until cracks among the shallow holes are communicated, and sufficient pressure relief is formed. Through the synergistic effect of shallow dense drilling pressure relief roadway protection and deep large-diameter reaming bearing energy release, a comprehensive structure integrating pressure relief, supporting and scour prevention and isolation is constructed, and the contradiction that roadway deformation and the pressure relief effect are difficult to consider at the same time in the prior art is effectively solved; and three effects of coal seam pressure relief, roadway support and rock burst prevention and control are integrated.

Description

Technical Field

[0001] This invention relates to the technical field of rockburst prevention and control methods, specifically to a rockburst prevention and control method and structure for mechanical borehole enlargement and unloading coupling in mines. Background Technology

[0002] In recent years, with the continuous increase in the depth and intensity of coal mining, coal and rock dynamic disasters such as mine tremors and rock bursts have become increasingly frequent, becoming one of the main hazards restricting safe production in deep mines. Rock bursts refer to the violent destructive phenomenon that occurs when the rock mass around the mine roadway or working face is released instantaneously due to strain energy. They are often accompanied by coal and rock mass ejection, loud noises, and strong air waves, and are extremely destructive. Therefore, designing effective rock burst prevention methods is crucial for achieving safe and efficient mining of such coal seams.

[0003] Using methods such as borehole decompression, hydraulic fracturing, and blasting to shift the peak bearing pressure of the coal seam to deeper layers is one of the main technical approaches for preventing rockbursts. Among these, borehole decompression is widely used due to its advantages of simple operation, low cost, and wide applicability. However, in actual engineering projects, multiple rounds of borehole decompression implemented to achieve the purpose of decompression and rockburst prevention can lead to a sharp increase in the width of the plastic zone of the coal seam next to the roadway, causing coal face anchorage failure and resulting in severe roadway deformation. Based on conventional drilling for pressure relief, patent CN119777872A proposes a mechanical borehole enlargement method for pressure relief in the elastic zone of coal seam: drilling a small-diameter hole in the plastic zone to penetrate into the elastic zone, and then activating the enlargement device to perform large-diameter enlargement in the peak zone of coal seam support pressure. Patent CN117390734A proposes a variable-diameter enlargement method for pressure relief parameter design: using small-diameter drilling to protect the support system in the roadway anchoring zone, and using large-diameter drilling for pressure relief in the coal seam stress concentration zone, aiming to effectively release coal stress while avoiding damage to the surrounding rock of the roadway. However, the above-mentioned mechanical borehole enlargement methods all enlarge the borehole from the peak zone of coal seam support pressure. Since this peak zone is usually located at the elastic-plastic boundary of the coal seam, the existing mechanical borehole enlargement method will still expand the width of the plastic zone of the coal seam next to the roadway, weakening the anchoring effect of the coal side anchor bolts. In addition, sufficient pressure relief in the enlarged area will significantly reduce the supporting effect of the coal seam on the roof in this area, leading to significant roof subsidence, inducing problems such as side heave and severe roadway deformation.

[0004] In summary, while existing mechanical borehole enlargement methods avoid the need for multiple rounds of pressure relief drilling during conventional borehole pressure relief, they still increase the width of the plastic zone in the coal seam next to the roadway, leading to coal side anchorage failure and severe roadway deformation. The contradiction between coal seam pressure relief and roadway support has not yet been resolved. Summary of the Invention

[0005] The purpose of this invention is to provide a method for preventing rockburst in mines by mechanically expanding the borehole and unloading the support, as well as a rockburst prevention structure, in order to solve the problems mentioned in the background art.

[0006] To achieve the above objectives, the present invention provides the following technical solution:

[0007] This application provides a mechanical borehole enlargement and unloading coupling method for preventing rockbursts in mines, including the following steps:

[0008] S1. Collect on-site geological data, determine the impact hazard of the coal seam based on the drill cuttings method, and determine the construction roadway;

[0009] S2. Calculate the construction parameters, and construct a single-layer conventional pressure relief borehole in the coal wall of the roadway according to the construction parameters; the conventional pressure relief borehole includes a shallow section and a deep section;

[0010] S3. Using a reaming drill bit, mechanical reaming is performed on the deep section of the conventional pressure relief drill hole to form a deep hole, and an elastic coal seam of a certain width is made between adjacent deep holes.

[0011] Optionally, the method further includes:

[0012] S4. Calculate the drilling pressure relief parameters in the shallow hole area, and drill conventional pressure relief shallow holes between the deep holes according to the pressure relief parameters;

[0013] S5. Monitor the pressure relief shallow holes. If the coal body between the shallow holes is not damaged, continue to drill shallow holes between the already drilled shallow holes until the cracks in the coal body between the shallow holes extend and penetrate, thus completing the mechanical hole enlargement and unloading-support coupling anti-scouring work of the coal seam.

[0014] Optionally, in step S2, the diameter D of the shallow hole section... q1 and the diameter D of the deep hole section s Determined based on the uniaxial compressive strength of the coal:

[0015] When the uniaxial compressive strength of the coal is less than 15 MPa, D q1 It is 150mm, D s It is 450mm;

[0016] When the uniaxial compressive strength of the coal body is 15-30 MPa, D q1 It is 200mm, D s It is 500mm;

[0017] When the uniaxial compressive strength of the coal body is greater than 30 MPa, D q1 It is 250mm, D s It is 550mm.

[0018] Optionally, in step S2, the length L1 of the shallow hole section is calculated using the following formula:

[0019]

[0020] Where C represents the coal cohesion in the borehole region. Let P be the friction angle within the coal seam in the borehole area, a be the roadway width, b be the roadway height, and P be the friction angle within the coal seam. i P0 represents the support pressure within the borehole area of ​​the tunnel, and P0 represents the original rock stress at the depth of the tunnel. All of these parameters are determined from the field geological data collected in step S1. The length L of the anchor bolt used for roadside anchoring is determined based on the anchor bolt dimensions used on-site and can be specified according to the on-site roadway excavation operation procedures. " is the floor operator.

[0021] Optionally, in step S2, the hole spacing B1 of the deep hole section is determined based on the length L1 of the shallow hole section, and the calculation formula is as follows:

[0022]

[0023] When 0m max When the depth is less than 2m, the hole spacing B1 in the deep hole section is determined to be 1.5m;

[0024] When B max When the depth is greater than 2m, the hole spacing B1 in the deep hole section is determined to be 2m.

[0025] Optionally, in step S2, the length L2 of the deep hole section is calculated using the following formula:

[0026]

[0027] The shallow hole length L1 and the deep hole spacing B1 are determined by step S2; the coal seam mining height H 采 Determined from the field geological data collected in step S1, " " is the floor operator.

[0028] Optionally, in step S4, the diameter D of the additional conventional pressure relief shallow hole is... q2 Through formula: D q2 =D q1 Sure,

[0029] Where D q1 The diameter of the shallow hole section mentioned in step S2.

[0030] Optionally, in step S4, the hole spacing B2 between the newly drilled shallow holes is determined by the formula:

[0031]

[0032] The hole spacing B3 between the shallow hole drilled and the deep hole after enlargement is determined by the formula:

[0033]

[0034] Where the diameter of the deep hole is D​s It is calculated from step S2.

[0035] Optionally, in step S4, the length L3 of the additional conventional pressure relief shallow hole is determined by the formula L3=L1+3;

[0036] Where L1 is the length of the shallow hole section described in step S2.

[0037] Secondly, this application provides a rockburst prevention structure, which is constructed in the coal seam of a rockburst mine using the above-mentioned method, and includes:

[0038] The pressure relief and roadway surrounding rock control zone is formed by conventional pressure relief boreholes;

[0039] The load-bearing and directional energy release zone is formed by mechanically enlarged holes;

[0040] An anti-scouring isolation zone located between the shallow and deep regions and consisting of an unexpanded elastic coal seam.

[0041] The present invention adopts the above technical solution and has the following significant advantages compared with the prior art:

[0042] To address the problems of large roof subsidence and unsatisfactory coal seam depressurization effects in existing borehole depressurization technologies, this invention achieves effective depressurization of coal seams in shallow borehole areas through the synergistic effect of single-layer dense conventional depressurization boreholes in shallow borehole areas and large-diameter, wide-spacing mechanical enlargement boreholes in deep borehole areas. Specifically, the single-layer dense conventional boreholes in shallow borehole areas provide ample deformation compensation space for the coal seam; while the elastic coal seam between the large-diameter, wide-spacing boreholes in deep borehole areas provides strong support for the roof, significantly suppressing roof subsidence. The synergy of these two methods reduces roof subsidence in shallow borehole areas while ensuring deformation compensation space for the coal seam in this region, thus guaranteeing effective depressurization.

[0043] To address the problems of expanded plastic zone in coal seams, coal face anchoring failure, and severe roadway deformation in existing technologies, this invention stabilizes and controls the surrounding rock of the roadway by implementing single-layer dense conventional pressure-relief drilling in shallow borehole areas. Specifically, the diameter of the shallow boreholes is much smaller than the roadway height, and the width of the plastic layer formed by the single-layer dense shallow boreholes is also much smaller than the roadway height, limiting disturbance to the surrounding rock. This maintains the integrity of the coal seam in the anchoring zone to ensure the anchoring effect of the anchor bolts, while effectively reducing the load on the surrounding rock through coal seam pressure relief, thus inhibiting the expansion of the plastic zone. These two aspects work synergistically to ensure the anchoring effect of the coal face and control the plastic zone of the surrounding rock, effectively solving the roadway support problem during coal seam pressure relief.

[0044] To address the limitation of existing technologies in terms of their single-layer anti-rockburst function, this invention integrates single-layer dense conventional pressure-relief drilling in shallow boreholes with large-diameter, wide-spacing mechanical enlargement in deep boreholes, constructing a multi-functional synergistic anti-rockburst system of "pressure relief in shallow boreholes - directional energy release in deep boreholes - elastic pressure-relief isolation." Specifically: effective pressure relief in shallow boreholes causes the peak coal seam support pressure to shift to the deep boreholes; the structurally weak surface formed by mechanical enlargement in deep boreholes provides energy release space for the coal seam strain energy in this area and changes the energy release direction (from the roadway to the deep borehole); the coal seam in the shallow pressure-relief elastic zone blocks the transmission of coal and rock kinetic energy from the deep boreholes to the roadway, forming an anti-rockburst isolation structure between the peak support pressure area and the roadway. The three elements work together to transfer the peak support pressure to the deep boreholes, provide instantaneous energy release space and change the energy release direction, while simultaneously constructing an elastic pressure-relief anti-rockburst isolation structure, achieving effective prevention and control of rockbursts.

[0045] In summary, this invention utilizes the synergistic effect of shallow, single-layer, dense conventional pressure relief boreholes and deep, large-diameter, wide-spacing mechanical enlargement boreholes to construct a multifunctional integrated structure that combines the functions of "pressure relief in shallow borehole areas - roadway protection - anti-rockburst isolation" with "bearing capacity in deep borehole areas - directional energy release," thereby achieving the triple technical objectives of coal seam pressure relief, roadway support, and rockburst prevention. Attached Figure Description

[0046] Figure 1 This is a schematic diagram of the drilling construction parameters of this invention;

[0047] Figure 2 This is a schematic diagram showing the distribution of the plastic zone within the xoy section at the borehole height after shallow drilling construction according to the present invention.

[0048] Figure 3 This is a schematic diagram showing the distribution of the plastic zone within the xoy section at the borehole height after mechanical reaming according to the present invention.

[0049] Figure 4 This is a schematic diagram showing the distribution of the plastic zone within the xoy section between holes after shallow drilling of the supplementary drilling section in this invention;

[0050] Figure 5 This is a schematic diagram showing the distribution of the plastic zone within the xoz section between holes after the drilling of this invention is completed;

[0051] Figure 6 This is a schematic diagram showing the distribution of the plastic zone within the xoz section of the hole after drilling is completed according to the present invention.

[0052] Figure 7 This is a diagram showing the vertical stress distribution of the coal seam 0.5m above the deep hole in an embodiment of the present invention.

[0053] Figure 8 This is a diagram showing the vertical stress distribution of the coal seam below the roof in an embodiment of the present invention.

[0054] Figure 9This is a construction flowchart of the rockburst mine mechanical borehole expansion and unloading-support coupling anti-rockburst method of the present invention. Detailed Implementation

[0055] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0056] Furthermore, elements in this invention are referred to as being "fixed to" or "set on" another element, which may be directly on the other element or may also include an intervening element. When an element is considered to be "connected" to another element, it may be directly connected to the other element or may also include an intervening element. The terms "vertical," "horizontal," "left," "right," and similar expressions used herein are for illustrative purposes only and do not represent the only possible implementations.

[0057] Example 1

[0058] The following detailed explanation of the proposed solution is illustrated with specific examples:

[0059] Please see Figures 1-9 This invention provides a mechanical enlargement and support coupling method for preventing rockburst in mines. This method introduces mechanical enlargement and supplementary drilling techniques on the basis of conventional pressure relief drilling, thereby achieving the technical goal of combining coal seam pressure relief, roadway support, and rockburst prevention in one process.

[0060] The implementation process of this invention is verified by combining on-site construction parameters and numerical simulation results. The following description uses a mine working face as an example. In this embodiment, a mine working face roadway is selected as the construction roadway. The roadway is 4.2m wide, 4.2m high, and approximately 650m deep. The coal seam thickness is 4.5m, and both the roof and floor are sandy mudstone. The roadway excavation direction is arranged along the coal seam, and the working face length is 300m. The solid coal sidewall beside the roadway is the area for this borehole enlargement and pressure relief, and the other side is a coal pillar sidewall.

[0061] Step S1. Collect on-site geological data and determine the working face construction roadway using the drill cuttings method. The geological data includes the following parameters: roadway width, height, depth, coal cohesion in the borehole area, internal friction angle, etc.

[0062] Step 101. Determine the working face construction roadway using the drill cuttings method. The borehole locations are the solid coal face of the mining roadway, with a monitoring area of ​​no less than 100m. The borehole spacing is 10m-30m, and at least 3 boreholes should be monitored on each side at a time. The monitoring interval is 1-3 days, and specific monitoring parameters can be adjusted according to the on-site construction situation. In this implementation case, the monitoring area is 150m, the borehole spacing is 15m, and the monitoring interval is 2 days. Step 102. Determine whether the roadway has an impact hazard based on the coal dust rate and dynamic effect indicators monitored during drilling. When the monitored coal dust rate index reaches the corresponding level, or when dynamic phenomena such as stuck drill, drill suction, drill top, abnormal noise, or in-hole impact occur during drilling, the coal body in the roadway is judged to have an impact hazard; otherwise, it is not. In this implementation case, abnormal noise and in-hole impact were observed in the monitored roadway, indicating an impact hazard in the roadway.

[0063] It should be noted that the coal dust rate index is used to determine the impact hazard at the work site by combining the borehole depth-to-height ratio (a) and the coal dust rate index (b). Specifically, the determination method is as follows: when the measured borehole depth-to-height ratio a < 1.5, if the coal dust rate index b ≥ 1.5, then an impact hazard is identified; when the borehole depth-to-height ratio a is between 1.5 and 3, the critical value of the coal dust rate index is 2; when the borehole depth-to-height ratio a > 3, the corresponding critical value of the coal dust rate index increases to 3.

[0064] Step 103. Based on the judgment results in Step 102, collect the corresponding roadway geological data, determine the roadway width as 4.2m, height as 4.2m, burial depth as approximately 650m, anchor bolt length as 2m, and the distance from the peak of the support pressure to the solid coal sidewall as 2.1m. The cohesion of the coal body is φ=34°, and the internal friction angle is C=5.2MPa. At the same time, the uniaxial compressive strength of the coal sample is calculated and determined to be 19MPa according to formula (1).

[0065] (1)

[0066] Step S2. Calculate the construction parameters and complete the conventional pressure relief drilling based on the obtained parameters.

[0067] The detailed construction steps are as follows:

[0068] Step 201. Determine the borehole height H: To facilitate on-site construction, the borehole height H should be 1m. During drilling, the borehole slope should be consistent with the roadway slope and perpendicular to the coal face.

[0069] Step 202. Determine the shallow hole diameter D q1 When the uniaxial compressive strength of the coal body calculated in step S1 is below 15 MPa, determine D. q1For a diameter of 150mm and a pressure of 15-30MPa, determine D. q1 For a thickness of 200mm; when the pressure is above 30MPa, determine D. q1 It is 250mm;

[0070] Step 203. Determine the deep hole diameter D s When the uniaxial compressive strength of the coal body calculated in step S1 is below 15 MPa, determine D. s For a diameter of 450mm and a pressure of 15-30MPa, determine D. s For a diameter of 500 mm; when the pressure is above 30 MPa, determine D. s It is 550mm.

[0071] Step 204. Determine the shallow hole length L1: The shallow hole length L1 can be calculated using the following formula (2).

[0072] (2)

[0073] In the formula: C is the cohesion of the coal in the borehole area, is the internal friction angle of the coal in the borehole area, a is the roadway width, b is the roadway height, and P is the internal friction angle of the coal in the borehole area. i P0 represents the support pressure within the borehole area of ​​the tunnel, and P0 represents the original rock stress at the depth of the tunnel. These parameters are determined by the field geological data collected in step S1. The length L of the anchor bolt used for anchoring along the tunnel is determined based on the size of the anchor bolt used on site. " is the floor operator.

[0074] Step 205. Determine the deep hole spacing B1: Based on the previously determined shallow hole length L1, the deep hole spacing B1 should satisfy formula (3). When 0m max When the depth is less than 2m, the deep hole spacing B1 is determined to be 1.5m; when B max When the depth is greater than 2m, the deep hole spacing B1 is determined to be 2m.

[0075] (3)

[0076] Step 206. Determine the total borehole length L: When the coal seam thickness is less than 3.5m, the total borehole length is determined to be 15m; when it is 3.5m-8m, the total borehole length is determined to be 20m; when it is greater than 8m, the total borehole length is determined to be 25m. At the same time, the length of the deep hole should not be less than 5 times the spacing between deep holes.

[0077] Step 207. Determine the deep hole length L2: The deep hole length L2 can be calculated using the following formula (4).

[0078] (4)

[0079] ​In the formula: the shallow hole length L1 is determined by step 204, and the deep hole spacing B1 is determined by step 205. The coal seam mining height H... 采 Determined from the field geological data collected in step S1, " " is the floor operator.

[0080] Step 208. The construction parameters for this example are calculated and determined through the above steps as follows: Shallow hole diameter D q1 =0.2m; Deep hole diameter D s =0.5m; deep hole spacing B1=2m; shallow hole length L1=6m; deep hole length L2=14m; drilling height H=1m. Complete conventional pressure relief drilling based on the above parameters.

[0081] Step 209. Complete conventional pressure relief drilling according to the parameters described in step 208.

[0082] Step S3. According to the parameters described in step 208: shallow hole length L1 = 6m, deep hole length L2 = 14m, deep hole diameter D s =0.5m, the deep hole section is mechanically enlarged using a reaming drill bit to form a large-diameter pressure relief space.

[0083] In one specific embodiment, the method provided in this application further includes:

[0084] Step S4. Calculate the construction parameters required for additional drilling, and drill conventional pressure relief holes between the enlarged boreholes according to the obtained parameters.

[0085] Step 401: Determine the diameter D of the additional borehole. q2 Determined through calculation using a formula.

[0086] Step 402: Determine the hole spacing B2 between the shallow holes to be drilled: Calculate using the following formula (5):

[0087] (5)

[0088] In the formula: the diameter of the shallow hole, D q2 Determined by step 401.

[0089] Step 403. Determine the hole spacing B3 between the supplementary drilling and the enlarged deep hole: This can be calculated using the following formula (6):

[0090] (6)

[0091] In the formula: the diameter of the shallow hole, D q2 The deep hole diameter D is determined by step 401. s Determined by step 202.

[0092] Step 404. Determine the length L3 of the additional drilling section: Based on the parameters mentioned in step 208: the shallow hole length L1 = 6m, it is determined by the formula.

[0093] Step 405. Based on the calculated parameters: the spacing between the shallow holes to be drilled is B2 = 0.25m, the spacing between the shallow holes to be drilled and the deep holes to be enlarged is B3 = 0.44m, and the length of the shallow holes to be drilled is L3 = 9m. Then, conventional pressure relief drill holes are drilled between the enlarged drill holes.

[0094] Step S5. Monitor the shallow boreholes and observe whether the coal body between the shallow boreholes is damaged. If the coal body between the shallow boreholes is not damaged, drill additional boreholes between the shallow boreholes. The diameter of the additional boreholes should be the same as the diameter of the additional boreholes, and the height of the additional boreholes should be the same as the height of the conventional boreholes, until the cracks in the coal body between the shallow boreholes extend and penetrate. At this point, the entire mechanical borehole enlargement and unloading coupling anti-rockburst method for rockburst mines is completed.

[0095] Experimental Example

[0096] To verify the effectiveness of the method of this invention, a FLAC3D numerical model consistent with the field was established. The model's length, width, and height were 50m × 6m × 21.5m, and it adopted the same rock strata structure and physical and mechanical parameters as the field. The coal seam and surrounding rock were set in layers, the model boundaries were fixed, vertical support constraints were applied at the bottom, and a uniformly distributed load of 17MPa was applied at the top to simulate the pressure of the overlying strata. The drilling and reaming locations and dimensions were consistent with the calculated construction parameters.

[0097] Numerical calculations show that after combined pressure relief through borehole enlargement and additional drilling, the stress in the shallow coal seam next to the roadway is significantly reduced, with the peak stress decreasing from 29.1 MPa to 24.5 MPa, a reduction of approximately 17%. The coal stress curve changes from a single peak to a multi-peak shape, with the maximum peak located in the deep borehole region. The internal stress field of the coal seam is redistributed, the strain energy in the shallow coal seam is effectively released, and the peak stress gradually shifts from the shallow to the deep region, forming a relatively reasonable pressure relief structure. After construction, the vertical stress distribution of the coal seam 0.5 m above the deep borehole is shown in the figure below. Figure 7 As shown in the figure, the vertical stress distribution of the coal seam below the roof is as follows: Figure 8 As shown.

[0098] In summary, compared to existing coal seam pressure relief and rockburst prevention methods, the mechanical borehole enlargement-support coupling method for rockburst-prone mines proposed in this invention achieves the technical goal of combining coal seam pressure relief, roadway support, and rockburst prevention through a comprehensive structural design that integrates pressure relief in shallow borehole areas, roadway protection, and rockburst prevention, as well as load-bearing and directional energy release in deep borehole areas. This provides an efficient and feasible technical approach for the prevention and control of rockbursts in mines. This method is widely applicable to coal seam pressure relief and rockburst prevention areas in rockburst-prone mines, with broad application prospects, laying a solid foundation for safe mining and improving mine production efficiency.

[0099] Secondly, this application provides a rockburst prevention structure, which is constructed in the coal seam of a rockburst mine using the above-mentioned method, and includes:

[0100] The pressure relief and roadway surrounding rock control zone is formed by conventional pressure relief boreholes;

[0101] The load-bearing and directional energy release zone is formed by mechanically enlarged holes;

[0102] An anti-scouring isolation zone located between the shallow and deep regions and consisting of an unexpanded elastic coal seam.

[0103] The pressure relief and roadway surrounding rock control zone effectively releases the strain energy of the shallow coal seam and enhances the stability of the roadway surrounding rock through the arrangement of conventional pressure relief boreholes. The bearing and directional energy release zone utilizes the elastic coal seam between the boreholes after mechanical enlargement to bear the stress peak transferred from the deep part, and uses mechanical enlargement to form a large cavity to release impact energy in a directional manner. The anti-impact isolation zone, on the other hand, uses the dense characteristics of the un-enlarged elastic coal seam to isolate the impact energy from propagating to the shallow part, forming an energy buffer zone. This structure, through the synergistic effect of the three zones, ensures the integrated realization of coal seam pressure relief, roadway support, and impact prevention, significantly improving mine safety and production efficiency.

[0104] It will be apparent to those skilled in the art that the present invention is not limited to the details of the exemplary embodiments described above, and that the invention can be implemented in other specific forms without departing from its spirit or essential characteristics. Therefore, the embodiments should be considered in all respects as exemplary and non-limiting, and the scope of the invention is defined by the appended claims rather than the foregoing description. Thus, all variations falling within the meaning and scope of equivalents of the claims are intended to be included within the present invention. No reference numerals in the claims should be construed as limiting the scope of the claims.

[0105] Furthermore, it should be understood that although this specification describes embodiments, not every embodiment contains only one independent technical solution. This narrative style is merely for clarity. Those skilled in the art should consider the specification as a whole, and the technical solutions in each embodiment can also be appropriately combined to form other embodiments that can be understood by those skilled in the art.

Claims

1. A method for preventing rockburst in mines using mechanical borehole enlargement and unloading coupling, characterized in that, Includes the following steps: S1. Collect on-site geological data, determine the impact hazard of the coal seam based on the drill cuttings method, and determine the construction roadway; S2. Calculate the construction parameters, and construct a single-layer conventional pressure relief borehole in the coal wall of the roadway according to the construction parameters; the conventional pressure relief borehole includes a shallow section and a deep section; S3. Using a reaming drill bit, mechanical reaming is performed on the deep section of the conventional pressure relief drill hole to form a deep hole, and an elastic coal seam of a certain width is made between adjacent deep holes.

2. The method according to claim 1, characterized in that, The method further includes: S4. Calculate the drilling pressure relief parameters in the shallow hole area, and drill conventional pressure relief shallow holes between the deep holes according to the pressure relief parameters; S5. Monitor the pressure relief shallow holes. If the coal body between the shallow holes is not damaged, continue to drill shallow holes between the already drilled shallow holes until the cracks in the coal body between the shallow holes extend and penetrate, thus completing the mechanical hole enlargement and unloading-support coupling anti-scouring work of the coal seam.

3. The method according to claim 1, characterized in that, In step S2, the diameter D of the shallow hole section q1 and the diameter D of the deep hole section s Determined based on the uniaxial compressive strength of the coal: When the uniaxial compressive strength of the coal is less than 15 MPa, D q1 It is 150mm, D s It is 450mm; When the uniaxial compressive strength of the coal body is 15-30 MPa, D q1 It is 200mm, D s It is 500mm; When the uniaxial compressive strength of the coal body is greater than 30 MPa, D q1 It is 250mm, D s It is 550mm.

4. The method according to claim 1, characterized in that, In step S2, the length L1 of the shallow hole section is calculated using the following formula: ； Where C represents the coal cohesion in the borehole region. Let P be the friction angle within the coal seam in the borehole area, a be the roadway width, b be the roadway height, and P be the friction angle within the coal seam. i P0 is the support pressure within the borehole area of ​​the tunnel, and P0 is the original rock stress at the depth of the tunnel. All these parameters are determined from the field geological data collected in step S1. The length Lanchor of the anchor bolt used for roadside anchoring is determined based on the anchor bolt dimensions used on site, and can be determined according to the on-site roadway excavation operation procedures. " is the floor operator.

5. The method according to claim 4, characterized in that, In step S2, the hole spacing B1 of the deep hole section is determined based on the length L1 of the shallow hole section, and the calculation formula is as follows: ； When 0m max When the depth is less than 2m, the hole spacing B1 in the deep hole section is determined to be 1.5m;​ When B max When the depth is greater than 2m, the hole spacing B1 in the deep hole section is determined to be 2m.

6. The method according to claim 1, characterized in that, In step S2, the length L2 of the deep hole section is calculated using the following formula: ； The shallow hole length L1 and the deep hole spacing B1 are determined by step S2; the coal seam mining height H 采 Based on the field geological data collected in step S1, " " is the floor operator.

7. The method according to claim 1, characterized in that, In step S4, the diameter D of the newly drilled conventional pressure relief shallow hole is... q2 Through formula: D q2 =D q1 Sure, Where D q1 The diameter of the shallow hole section mentioned in step S2.

8. The method according to claim 7, characterized in that, In step S4, the hole spacing B2 between the newly drilled shallow holes is determined by the formula: ； The hole spacing B3 between the shallow hole drilled and the deep hole after enlargement is determined by the formula: ； Where the diameter of the deep hole is D s It is calculated from step S2.

9. The method according to claim 1, characterized in that, In step S4, the length L3 of the additional conventional pressure relief shallow hole is determined by the formula L3=L1+3; Where L1 is the length of the shallow hole section described in step S2.

10. A rockburst prevention structure, characterized in that, The structure is constructed in a coal seam of a rockburst mine using the method described in any one of claims 1 to 9, and comprises: The pressure relief and roadway surrounding rock control zone is formed by conventional pressure relief boreholes; The load-bearing and directional energy release zone is formed by mechanically enlarged holes; An anti-scouring isolation zone located between the shallow and deep regions and consisting of an unexpanded elastic coal seam.

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