A method for preventing and treating coal body spalling in deep rock burst roadway tunneling

Abstract

The application provides a deep rock burst roadway tunneling coal body spalling prevention method, and the prevention method comprises the following steps: collecting undisturbed coal samples in a roadway tunneling head area; determining a drilling hole diameter and a hole spacing through a comparative experiment; expanding a side of a roadway tunneling working face to form a drilling field; drilling multiple drilling holes in the drilling field along a tunneling direction of the roadway and forming multiple drilling hole groups; one drilling hole group comprises a static blasting hole and two free holes; performing a cutting seam guide treatment on the static blasting hole; plugging the static blasting hole after injecting a static blasting agent into the static blasting hole; monitoring crack propagation and energy release characteristics of the static blasting hole until the crack propagation and the energy release tend to be stable, and pressure relief is completed; after the pressure relief is completed, tunneling and anchor-mesh-cable combined support are implemented. The prevention method ensures that a side part coal body is stable during tunneling, improves a tunneling coal body spalling problem, and improves anchor-mesh-cable support effect and long-term stability of a roadway.

Description

Technical Field

[0001] This invention belongs to the field of coal mine roadway surrounding rock control technology, specifically, it relates to a method for preventing coal body spalling during deep rockburst roadway excavation. Background Technology

[0002] As coal mining extends to deeper levels, high ground stress and intense mining disturbances coexist, making rockbursts a major hazard for safe and efficient mining in deep coal mines. Rockbursts are a dynamic phenomenon caused by the sudden release of elastic strain energy in coal and rock masses under high stress, leading to severe deformation and damage of the surrounding rock in the roadway. Among these phenomena, coal spalling is the most typical form of surrounding rock failure during the excavation of roadways prone to rockbursts in deep mines.

[0003] In deep areas prone to rock bursts, even with high coal seam strength, well-developed joints and fissures can lead to frequent stress disturbances during tunnel excavation, which, combined with the original high ground stress, create a complex dynamic-static high-stress field. As a result, the coal seam accumulates a large amount of elastic energy. When the excavation disturbance exceeds its bearing capacity, the coal seam at the tunnel face and on both sides often experiences instantaneous and sudden fracturing and collapse, accompanied by noticeable coal blasting sounds—this is known as the coal face spalling phenomenon.

[0004] In related technologies, the prevention and control of rock bursts typically focuses on "post-event pressure relief," such as large-diameter drilling to relieve pressure on the rock face, deep-hole blasting, or hydraulic fracturing. However, rock bursts often occur instantly after the roadway is exposed, rendering these methods ineffective due to their delayed effect. Current mainstream "advanced pretreatment" technologies, such as drilling parallel axial boreholes on both sides in front of the roadway face for pre-fracture pressure relief, can alleviate rock bursts to some extent, but still suffer from limited fracturing range, insufficient pressure relief capacity, and high risk of rock bursts under the special high-stress environment of rock burst roadways. Summary of the Invention

[0005] The present invention aims to at least partially solve one of the technical problems in the related art.

[0006] Therefore, embodiments of the present invention propose a method for preventing coal seam spalling during deep rockburst roadway excavation. This method ensures the stability of the coal seam sidewalls during excavation, improves the problem of coal seam spalling, enhances the effect of anchor wire mesh support and the long-term stability of the roadway, eliminates the need for frequent repairs, and comprehensively guarantees construction safety and excavation efficiency.

[0007] The method for preventing coal seam spalling during deep rockburst tunnel excavation according to an embodiment of the present invention includes the following steps:

[0008] The original ground stress in the target roadway area was measured, and undisturbed coal samples were collected at the roadway excavation face. The borehole diameter and spacing were determined through indoor static explosion effect comparison experiments. Enlarge at least one side of the tunnel excavation face to form a drilling site; Within the drilling site, multiple boreholes are drilled along the advancing direction of the roadway, forming an axial borehole group for the surrounding rock of the sidewall. The borehole group includes multiple borehole sets, and each borehole set includes a static blasting hole and two free holes. The two free holes are symmetrically arranged on the upper and lower sides of the static blasting hole in the vertical direction. The diameters of the static blasting hole and the free holes, as well as the distance between the static blasting hole and the free holes, are determined based on indoor static blasting effect comparison experiments. The static blasting hole is subjected to slit guiding treatment. Slits are made on the upper and lower sides of the hole wall along the extension direction of the static blasting hole, and the extension direction of the slits is toward the extension direction of the free hole. After injecting detonating agent into the detonation hole, the detonation hole is sealed. Monitor the crack propagation and energy release characteristics in static explosion holes until crack propagation and energy release tend to stabilize, and complete the pressure relief; After depressurization, tunneling is carried out and anchor-mesh-cable combined support is implemented.

[0009] The method for preventing coal seam spalling in deep rockburst roadways according to this invention achieves a precise match between advanced directional stress relief and coal seam reinforcement in rockburst roadways through the synergistic innovation of static blast hole parameters determined by laboratory optimization, vertical slit guidance technology, free hole guidance unloading and water replenishment design, and polyurethane sealing technology. The arrangement of static blast holes and free holes, combined with vertical slit treatment, forcibly guides static blast cracks to develop vertically, avoiding disorderly expansion in high-stress fields. This precisely transfers concentrated stress in the roadway walls and rapidly releases the elastic energy accumulated in the coal, fundamentally curbing spalling. This ensures the stability of the coal seam walls during excavation, improves the problem of coal seam spalling, enhances the effectiveness of anchor wire mesh support and the long-term stability of the roadway, eliminates the need for frequent repairs, and comprehensively guarantees construction safety and excavation efficiency.

[0010] In some embodiments, the indoor static explosion effect comparison experiment further includes the following steps: The collected undisturbed coal samples were divided into two experimental groups, with each group containing multiple undisturbed coal samples. A single borehole was drilled at the center of the undisturbed coal sample in the first experimental group. Multiple first parallel specimens were prepared, with boreholes of different diameters forming a first control group. Based on the field stress measurement results, vertical and horizontal confining pressures were applied to the specimens. After the confining pressure was applied, a static explosive was injected into the boreholes, and the crack initiation time, propagation range, and fracturing effect were monitored to determine the borehole diameter. In the second experimental group, the undisturbed coal samples were equipped with a vertical single-row double-hole configuration. The upper hole was a free hole, and the lower hole was a static explosion hole. The diameter of the static explosion hole was the same as that determined in the first experimental group. Multiple second parallel samples were set up, with different distances between the holes of the second parallel samples, which served as a second comparison group. The same vertical and horizontal confining pressure as in the first experimental group was applied. After the confining pressure was applied, a static explosion agent was injected into the holes, and the crack initiation time, propagation range, and fracturing effect were monitored to determine the distance between the static explosion hole and the free hole.

[0011] In some embodiments, the diameter of the borehole is greater than or equal to 40 mm and less than or equal to 100 mm, and at least 3 parallel samples are provided for each borehole diameter; The hole spacing in the second parallel specimen is greater than 100 mm and less than or equal to 300 mm, and at least 3 parallel specimens are set for each group of spacing.

[0012] In some embodiments, the diameter of the free hole is smaller than the diameter of the static explosion hole.

[0013] In some embodiments, the distance between two adjacent sets of boreholes is greater than or equal to 300 mm and less than or equal to 500 mm.

[0014] In some embodiments, the depth of the slit is 1 / 3 to 1 / 2 of the radius of the static blast hole.

[0015] In some embodiments, the step of removing all rock powder and debris from the borehole is included before injecting the detonating agent into the detonation hole.

[0016] In some embodiments, injecting the detonating agent into the detonation hole further includes the following steps: The static blasting agent and water are rapidly mixed at a water-cement ratio of 0.3 to form a static blasting agent slurry. The mixing time for each stage is controlled at 1 to 3 minutes. After mixing, the slurry is immediately injected into the static blasting hole in stages using a grouting pump, stopping when the slurry is 1 to 1.5 meters away from the hole opening.

[0017] In some embodiments, the method for preventing coal seam spalling during deep rockburst tunnel excavation according to the present invention further includes the following steps: 30-60 minutes after the static detonator slurry is injected, the hydration reaction begins. When the acoustic emission monitoring equipment detects signs of microcrack expansion in the coal body around the static detonator hole, clean water is injected into all free holes at a pressure of 1-3 MPa for 10-20 minutes to replenish the water evaporated during the hydration reaction and ensure that the detonator reacts fully.

[0018] In some embodiments, sealing the static explosion hole includes the following steps: filling and sealing the static explosion hole with polyurethane material, with a sealing length of 1~1.5m. Attached Figure Description

[0019] Figure 1 This is a schematic diagram illustrating the implementation of the method for preventing coal seam spalling during deep rockburst tunnel excavation according to an embodiment of the present invention.

[0020] Figure 2 This is a schematic diagram of the cutting method in the deep rockburst roadway excavation coal seam spalling prevention method according to an embodiment of the present invention.

[0021] Figure 3 yes Figure 2 A cross-sectional view of the central cut.

[0022] Figure 4 This is a schematic diagram of crack propagation in the method for preventing coal seam spalling during deep rockburst tunnel excavation according to an embodiment of the present invention.

[0023] Figure 5 This is a schematic diagram showing the distribution of static blast holes and free holes in the method for preventing coal seam spalling during deep rockburst tunnel excavation according to an embodiment of the present invention.

[0024] Figure label: 100. Coal body, 200. Tunnel excavation face. 1. Static blast hole, 2. Free hole, 3. Cutting the seam. 4. Cracks. Detailed Implementation

[0025] Embodiments of the present invention are described in detail below, examples of which are illustrated in the accompanying drawings. The embodiments described below with reference to the accompanying drawings are exemplary and intended to explain the present invention, and should not be construed as limiting the present invention.

[0026] like Figures 1-5 As shown, the method for preventing coal seam spalling during deep rockburst tunnel excavation according to an embodiment of the present invention includes the following steps: The original ground stress in the target roadway area was measured, and undisturbed coal samples were collected at the roadway excavation face.

[0027] Understandably, geostress testing techniques (such as hydraulic fracturing and stress relief methods) can be used to determine the original geostress field in the tunnel excavation area, including the maximum principal stress, minimum principal stress, and direction.

[0028] Uncirculated coal samples were collected from the tunnel face area for laboratory testing to determine the mechanical parameters of the coal body (such as compressive strength, elastic modulus, and Poisson's ratio) and the characteristics of joint and fracture development. This was done to accurately understand the stress environment and mechanical properties of the coal body, providing data support for subsequent borehole parameter optimization. This also helps avoid the failure of stress relief schemes or excessive damage to the surrounding rock due to misjudgment of the stress field.

[0029] The borehole diameter and spacing were determined through indoor static explosion effect comparison experiments. It is understood that the collected coal samples can be divided into multiple groups for indoor static explosion effect comparison experiments. Within the same group, different borehole diameters can be used for comparison, or static explosion borehole 1 and free borehole 2 can be compared.

[0030] In other words, multiple sets of indoor static explosion effect comparison experiments can be used to determine the optimal aperture and aperture spacing, which facilitates crack 4-directional propagation and efficient pressure relief, avoiding excessive damage to the surrounding rock due to excessive aperture or insufficient pressure relief due to excessive aperture.

[0031] like Figure 5 As shown, at least one side of the tunnel excavation face 200 is widened to form a drilling site.

[0032] It is understandable that the tunnel face 200 is widened on one or both sides to form a dedicated fracturing drilling site for static blasting. Preferably, the widening width is 1.5~2.5 m to ensure sufficient operating space for drilling and to not affect subsequent tunneling.

[0033] Within the drilling site, multiple boreholes are drilled along the advancing direction of the tunnel, forming an axial borehole group for the surrounding rock of the sidewall. The borehole group includes multiple borehole sets. Each borehole set includes a static blast hole 1 and two free holes 2. The two free holes 2 are symmetrically arranged on the upper and lower sides of the static blast hole 1 in the vertical direction. The diameters of the static blast hole 1 and the free holes 2, as well as the distance between the static blast hole 1 and the free holes 2, are determined based on indoor static blast effect comparison experiments.

[0034] Understandably, a crawler-mounted drilling rig can be used to drill a group of holes along the tunnel axis. The diameter of the static blasting hole 1 can be determined by referring to the values ​​tested in the laboratory, and the distance between the static blasting hole 1 and the free hole 2 is also determined by the values ​​tested in the laboratory.

[0035] In other words, the borehole group forms an advanced pressure relief zone, releasing 100% of the elastic energy of the coal body ahead of the tunneling, and reducing the risk of dynamic disturbance-induced fracturing. Free hole 2 can serve as a stress transfer channel to guide the directional propagation of static explosion crack 4, avoiding disordered fracturing.

[0036] like Figures 1-4 As shown, the static blasting hole 1 is guided by a slit 3. Slits 3 are made on the upper and lower sides of the hole wall of the static blasting hole 1 along the extension direction of the static blasting hole 1, and the extension direction of the slit 3 is towards the extension direction of the free hole 2.

[0037] It is understandable that an abrasive jet injector can be used for the slit treatment 3. That is, the abrasive jet injector is inserted into the static explosion hole 1, and high-pressure abrasive jet technology is used to perform segmented vertical slit treatment 3 on the hole wall. The forced crack 4 extends along the direction of slit 3 (vertical). Slit 3 connects to the free hole 2 to form a pressure relief channel, avoiding uneven pressure relief or local stress concentration caused by the random extension of crack 4 in the high stress field.

[0038] After injecting the detonating agent into the detonation hole 1, the detonation hole 1 is sealed.

[0039] Understandably, polyurethane material with both high strength and good sealing properties can be selected. After being mixed evenly according to the product ratio, it can be injected into the filling and sealing section of the static explosion hole 1 to ensure that the hole opening is completely sealed, prevent the static explosion agent slurry from being lost or blowouts from occurring during the reaction process, and at the same time protect the integrity of the shallow surrounding rock of the roadway.

[0040] Monitor the propagation and energy release characteristics of crack 4 in static blast hole 1 until the propagation and energy release of crack 4 tend to stabilize, and complete the depressurization. After depressurization, tunneling is carried out and anchor-mesh-cable combined support is implemented.

[0041] Understandably, once the pressure relief is completed, the stress in the 100-ton coal seam of the side wall is fully released and the stability meets the requirements, normal tunneling can proceed. After tunneling, anchor-mesh-cable combined support is implemented promptly. Because the 100-ton coal seam of the side wall has undergone directional static blasting for pressure relief, and the shallow surrounding rock remains intact, the preload loss of the anchor bolts (cables) is significantly reduced, the support effect is greatly improved, and the roadway does not require frequent repairs.

[0042] Therefore, the method for preventing coal seam spalling in deep rockburst roadway excavation according to this invention achieves a precise match between advanced directional stress relief and the reinforcement of the coal seam 100 in rockburst roadways through the synergistic innovation of the parameters of the static blasting hole 1 determined by laboratory optimization, the vertical slit 3 guiding technology, the free hole 2 guiding unloading and water replenishment design, and the polyurethane sealing technology. The arrangement of the static blasting hole 1 and the free hole 2, combined with the vertical slit 3 treatment, forcibly guides the static blasting crack 4 to develop vertically and directionally, avoiding disorderly expansion in a high-stress field, accurately transferring the concentrated stress in the roadway sidewalls, and rapidly releasing the elastic energy accumulated in the coal seam 100, thus curbing spalling from the root. This ensures the stability of the coal seam 100 during excavation, improves the spalling problem of the excavated coal seam 100, enhances the anchor mesh cable support effect and the long-term stability of the roadway, eliminates the need for frequent repairs, and comprehensively guarantees construction safety and excavation efficiency.

[0043] In some embodiments, the indoor static explosion effect comparison experiment further includes the following steps: dividing the collected undisturbed coal samples into a first experimental group and a second experimental group, with each experimental group including multiple undisturbed coal samples.

[0044] Understandably, the collected undisturbed coal samples are divided into a first experimental group (single-hole experiment) and a second experimental group (dual-hole experiment), with each group containing at least 3-5 parallel samples to ensure data reliability. Preferably, CT scans or acoustic tests are performed on the coal samples before the experiment to record the initial fracture development. This eliminates the influence of individual sample differences on the experimental results and improves the scientific rigor of parameter determination. Comparison of parallel samples verifies the repeatability and universality of the experimental results.

[0045] A single borehole was set in the center of the undisturbed coal sample in the first experimental group, and multiple first parallel samples were set up. The boreholes of the first parallel samples were of different diameters and served as the first control group. According to the on-site in-situ stress measurement results, vertical and horizontal confining pressures were applied to the samples. After the confining pressure was completed, a static explosive was injected into the borehole, and the crack initiation time, propagation range and crushing effect were monitored to determine the borehole diameter.

[0046] Understandably, multiple first parallel specimens can be arranged with a borehole diameter difference of 5 mm or 10 mm. Vertical and horizontal confining pressures are applied to the specimens on a triaxial press, with the stress level consistent with the measured in-situ stress, and the front and rear ends of the specimens are rigidly constrained. A static blasting agent slurry is prepared with a water-cement ratio of 0.3 and rapidly injected into the borehole. Strain monitoring, digital image correlation (DIC), acoustic emission, and CT scanning techniques are used to monitor the crack initiation time, propagation range, and fracturing effect of crack 4 in real time. The optimal diameter of the static blasting hole (hereinafter referred to as "static blasting hole 1") is determined using the uniformity of crack propagation, the effective fracturing range compliance rate, and the crack initiation efficiency as evaluation indicators.

[0047] In the second experimental group, the original coal sample was set with a vertical single row of double holes. The upper hole was a free hole 2, and the lower hole was a static explosion hole 1. The diameter of the static explosion hole 1 was the same as that determined in the first experimental group. Multiple second parallel samples were set up. The distance between the holes of the second parallel samples was different, and they served as the second comparison group. The same vertical and horizontal confining pressure as in the first experimental group was applied. After the confining pressure was applied, the static explosion agent was injected into the hole, and the crack initiation time, propagation range and crushing effect of crack 4 were monitored to determine the distance between the static explosion hole 1 and the free hole 2.

[0048] Multiple second parallel specimens were used, with the hole spacing varying by 50 mm or 100 mm. The diameter of the static explosion hole 1 was determined using the optimal value from the first parallel specimen experiment. The same stress conditions as the first parallel specimen were applied, and static explosion agent slurry was injected only into the static explosion hole 1. Using the same monitoring methods, the optimal spacing between the static explosion hole 1 and the free hole 2 was determined based on the directional propagation effect of crack 4, the fracture range, and the energy release efficiency.

[0049] Preferably, the borehole diameter is greater than or equal to 40 mm and less than or equal to 100 mm, and at least 3 parallel specimens are set for each borehole diameter; the hole spacing in the second parallel specimen is greater than 100 mm and less than or equal to 300 mm, and at least 3 parallel specimens are set for each spacing.

[0050] Understandably, the borehole diameters were set to 40 mm, 50 mm, 60 mm, 70 mm, 80 mm, 90 mm, and 100 mm, with three parallel specimens for each diameter. The borehole spacings were 100 mm, 150 mm, 200 mm, 250 mm, and 300 mm, with three parallel specimens for each spacing.

[0051] In some embodiments, such as Figure 5 As shown, the diameter of the free hole 2 is smaller than the diameter of the static explosion hole 1.

[0052] Understandably, the static detonation hole 1 generates enormous expansion pressure through the chemical reaction of the static detonating agent, using this pressure to fracture the coal body 100 and drive the propagation of crack 4. Therefore, it needs sufficient space to fill with a adequate amount of static detonating agent to ensure sufficient energy to complete the pressure relief operation. The free hole 2 serves to guide the directional propagation of the static detonation crack 4. When crack 4 originates from the static detonation hole 1 and propagates towards the free hole 2, a stress concentration zone forms on the wall of the free hole 2.

[0053] In other words, if the diameter of the free hole 2 is similar to or even larger than that of the static explosion hole 1, the resulting low-pressure zone and stress relief surface will be too large. This may cause the expansion pressure energy generated by the static explosion hole 1 to disperse prematurely towards the free hole 2 instead of forming a concentrated splitting effect, thereby reducing the propagation efficiency and depth of crack 4 and resulting in insufficient pressure relief.

[0054] Preferably, the diameter of the free hole 2 is half the diameter of the static explosion hole 1.

[0055] Preferably, the distance between two adjacent borehole groups is greater than or equal to 300 mm and less than or equal to 500 mm.

[0056] Understandably, different hole spacings can be set according to actual working conditions to ensure that there are sufficient rock pillars between borehole groups to maintain the basic bearing capacity of the surrounding rock, avoiding excessive fragmentation and resource waste. It also ensures that the decompression radius of a single group is sufficient to cover adjacent groups, thus forming a seamless, integrated pre-decompression zone and eliminating the risk of residual stress.

[0057] Preferably, the depth of the slit 3 is 1 / 3 to 1 / 2 of the radius of the static blast hole 1.

[0058] It is understandable that, such as Figure 3 and Figure 4As shown, a shallow cut 3, if its depth is less than 1 / 3 of the radius of the static detonation hole 1, provides an insufficiently clear guiding path. Even if crack 4 originates from here, it may deviate from its direction in the early stages of propagation, becoming subject to the stress field or natural fissures again, and failing to effectively connect to the free hole 2, ultimately leading to uneven pressure relief. Conversely, if the cut 3 is too deep, premature failure of the hole wall will cause leakage of the static detonation agent slurry, preventing effective pressure accumulation and causing the entire pressure relief operation to fail.

[0059] In some embodiments, the step of removing all rock powder and debris from the borehole is included before injecting the detonating agent into the detonation hole 1.

[0060] Understandably, the pressure relief principle of the static detonator is that the expansion pressure generated by its chemical reaction acts uniformly on the coal body 100 through the borehole wall. If rock powder and debris remain in the borehole, when the static detonator begins to expand, most of the energy will be used first to compact these rock powder and debris, rather than acting directly on the borehole wall. This leads to a huge waste of energy, significantly reducing the effective pressure actually transmitted to the coal body 100. This energy attenuation may cause the static detonation pressure to fail to reach the fracturing threshold of the coal body 100, or the fracturing time to be delayed or insufficient, ultimately resulting in insufficient pressure relief and rendering the entire prevention and control measure ineffective.

[0061] In some embodiments, injecting the detonating agent into the detonation hole 1 further includes the following steps: rapidly stirring the detonating agent and water at a water-cement ratio of 0.3 to form a detonating agent slurry, with the single-stage stirring time controlled at 1 to 3 minutes, and immediately injecting the slurry into the detonation hole 1 in a segmented injection manner using a grouting pump after stirring, stopping when the slurry is 1 to 1.5 m away from the hole opening.

[0062] It is understandable that, such as Figure 1 As shown, the hydration reaction of the static detonator (static fracturing agent) is the fundamental cause of the expansion pressure. The hydration reaction begins the moment the static detonator comes into contact with water. If the stirring time is too long, the slurry may begin to set in the pumping pipeline or grouting tank, leading to blockage, uneven filling, or even failure.

[0063] For long, vertical or inclined boreholes, if the grout is injected all at once from the borehole opening, the grout is prone to solid particle sedimentation under gravity, resulting in an uneven distribution of grout with a dense bottom and a thin opening. Using segmented injection (e.g., gradually pulling out the grouting pipe and injecting once after each segment) ensures a more uniform distribution of grout along the entire length of the borehole, resulting in more balanced stress on the borehole wall and more coordinated crack propagation.

[0064] like Figure 1 and Figure 2 As shown, the grout is poured to a point 1-1.5m from the borehole opening to facilitate subsequent sealing.

[0065] In some embodiments, the method for preventing coal body spalling during deep rockburst tunneling according to the present invention further includes the following steps: 30-60 minutes after the injection of the static explosive agent slurry, the hydration reaction begins. When the acoustic emission monitoring device detects signs of microcrack expansion in the coal body 100 around the static explosive hole 1, clean water is injected into all free holes 2 at a pressure of 1-3 MPa for 10-20 minutes to replenish the water evaporated by the hydration reaction and ensure that the fracturing agent reacts fully.

[0066] Understandably, the hydration reaction of the detonating agent is exothermic. Within the sealed borehole, the temperature rises significantly, leading to water evaporation. Water is essential for the reaction, and its loss can slow down or even halt it. The period from 30 to 60 minutes after the reaction begins is the most intense period of exothermic hydration and the fastest water evaporation. During this time, the detection of microcrack propagation using acoustic emission (AE) equipment confirms that the reaction is ongoing and effective, thus determining the window for water replenishment.

[0067] Pressurized water injection into free hole 2 allows the water to flow precisely through these tiny crack channels into the area where the reaction is taking place, directly replenishing the evaporated water. This ensures that the detonating agent continues to react fully, maintaining and increasing the expansion pressure, allowing crack 4 to extend further and more completely.

[0068] In some embodiments, such as Figure 1 and Figure 2 As shown, sealing the static explosion hole 1 includes the following steps: filling and sealing the static explosion hole 1 with polyurethane material, with a sealing length of 1~1.5m.

[0069] Understandably, polyurethane expands in volume after the reaction, allowing it to tightly adhere to the irregular pore walls, forming a reliable frictional lock and adhesive seal to prevent leakage. Furthermore, using polyurethane to fill and seal the static explosion pore 1 can quickly generate sufficiently high mechanical strength to withstand the enormous expansion pressure generated by the static explosion agent.

[0070] Preferably, the sealing length is 1 to 1.5 m, which is an engineering safety length that can reliably resist expansion pressure.

[0071] In the description of this invention, it should be understood that the terms "center," "longitudinal," "lateral," "length," "width," "thickness," "upper," "lower," "front," "rear," "left," "right," "vertical," "horizontal," "top," "bottom," "inner," "outer," "clockwise," "counterclockwise," "axial," "radial," and "circumferential" indicate the orientation or positional relationship based on the orientation or positional relationship shown in the accompanying drawings. They are used only for the convenience of describing this invention and simplifying the description, and are not intended to indicate or imply that the device or element referred to must have a specific orientation, or be constructed and operated in a specific orientation. Therefore, they should not be construed as limitations on this invention.

[0072] Furthermore, the terms "first" and "second" are used for descriptive purposes only and should not be construed as indicating or implying relative importance or implicitly specifying the number of technical features indicated. Thus, a feature defined as "first" or "second" may explicitly or implicitly include at least one of that feature. In the description of this invention, "a plurality of" means at least two, such as two, three, etc., unless otherwise explicitly specified.

[0073] In this invention, unless otherwise explicitly specified and limited, the terms "installation," "connection," "linking," and "fixing," etc., should be interpreted broadly. For example, they can refer to a fixed connection, a detachable connection, or an integral part; they can refer to a mechanical connection, an electrical connection, or a connection that allows communication between them; they can refer to a direct connection or an indirect connection through an intermediate medium; they can refer to the internal communication of two components or the interaction between two components, unless otherwise explicitly limited. Those skilled in the art can understand the specific meaning of the above terms in this invention according to the specific circumstances.

[0074] In this invention, unless otherwise explicitly specified and limited, "above" or "below" the second feature can mean that the first feature is in direct contact with the second feature, or that the first feature is in indirect contact with the second feature through an intermediate medium. Furthermore, "above," "over," and "on top" of the second feature can mean that the first feature is directly above or diagonally above the second feature, or simply that the first feature is at a higher horizontal level than the second feature. "Below," "below," and "under" the second feature can mean that the first feature is directly below or diagonally below the second feature, or simply that the first feature is at a lower horizontal level than the second feature.

[0075] In this invention, the terms "one embodiment," "some embodiments," "example," "specific example," or "some examples," etc., refer to a specific feature, structure, material, or characteristic described in connection with that embodiment or example, which is included in at least one embodiment or example of the invention. In this specification, the illustrative expressions of the above terms do not necessarily refer to the same embodiment or example. Furthermore, the specific features, structures, materials, or characteristics described may be combined in any suitable manner in one or more embodiments or examples. Moreover, without contradiction, those skilled in the art can combine and integrate the different embodiments or examples described in this specification, as well as the features of different embodiments or examples.

[0076] Although embodiments of the present invention have been shown and described above, it is understood that the above embodiments are exemplary and should not be construed as limiting the present invention. Those skilled in the art can make changes, modifications, substitutions and variations to the above embodiments within the scope of the present invention.

Claims

1. A method for preventing coal seam spalling during deep rockburst roadway excavation, characterized in that, Includes the following steps: The original ground stress in the target roadway area was measured, and undisturbed coal samples were collected at the roadway excavation face. The borehole diameter and spacing were determined through indoor static explosion effect comparison experiments. Enlarge at least one side of the tunnel excavation face to form a drilling site; Within the drilling site, multiple boreholes are drilled along the advancing direction of the roadway, forming an axial borehole group for the surrounding rock of the sidewall. The borehole group includes multiple borehole sets, and each borehole set includes a static blasting hole and two free holes. The two free holes are symmetrically arranged on the upper and lower sides of the static blasting hole in the vertical direction. The diameters of the static blasting hole and the free holes, as well as the distance between the static blasting hole and the free holes, are determined based on indoor static blasting effect comparison experiments. The static blasting hole is subjected to slit guiding treatment. Slits are made on the upper and lower sides of the hole wall along the extension direction of the static blasting hole, and the extension direction of the slits is toward the extension direction of the free hole. After injecting detonating agent into the detonation hole, the detonation hole is sealed. Monitor the crack propagation and energy release characteristics in static explosion holes until crack propagation and energy release tend to stabilize, and complete the pressure relief; After depressurization, tunneling is carried out and anchor-mesh-cable combined support is implemented.

2. The method for preventing coal seam spalling during deep rockburst tunnel excavation according to claim 1, characterized in that, The indoor static explosion effect comparison experiment also includes the following steps: The collected undisturbed coal samples were divided into two experimental groups, with each group containing multiple undisturbed coal samples. A single borehole was drilled at the center of the undisturbed coal sample in the first experimental group. Multiple first parallel specimens were prepared, with boreholes of different diameters forming a first control group. Based on the field stress measurement results, vertical and horizontal confining pressures were applied to the specimens. After the confining pressure was applied, a static explosive was injected into the boreholes, and the crack initiation time, propagation range, and fracturing effect were monitored to determine the borehole diameter. In the second experimental group, the undisturbed coal samples were equipped with a vertical single-row double-hole configuration. The upper hole was a free hole, and the lower hole was a static explosion hole. The diameter of the static explosion hole was the same as that determined in the first experimental group. Multiple second parallel samples were set up, with different distances between the holes of the second parallel samples, which served as a second comparison group. The same vertical and horizontal confining pressure as in the first experimental group was applied. After the confining pressure was applied, a static explosion agent was injected into the holes, and the crack initiation time, propagation range, and fracturing effect were monitored to determine the distance between the static explosion hole and the free hole.

3. The method for preventing coal seam spalling during deep rockburst tunnel excavation according to claim 2, characterized in that, The diameter of the borehole is greater than or equal to 40 mm and less than or equal to 100 mm, and at least 3 parallel samples are set for each borehole diameter; The hole spacing in the second parallel specimen is greater than 100 mm and less than or equal to 300 mm, and at least 3 parallel specimens are set for each group of spacing.

4. The method for preventing coal seam spalling during deep rockburst tunnel excavation according to claim 2, characterized in that, The diameter of the free hole is smaller than the diameter of the static explosion hole.

5. The method for preventing coal seam spalling during deep rockburst tunnel excavation according to claim 1, characterized in that, The distance between two adjacent borehole groups is greater than or equal to 300 mm and less than or equal to 500 mm.

6. The method for preventing coal seam spalling during deep rockburst tunnel excavation according to claim 1, characterized in that, The depth of the cut is 1 / 3 to 1 / 2 of the radius of the static blast hole.

7. The method for preventing coal seam spalling during deep rockburst tunnel excavation according to claim 1, characterized in that, Before injecting the detonating agent into the detonation hole, the following steps are also included: removing all rock powder and debris from the borehole.

8. The method for preventing coal seam spalling during deep rockburst tunnel excavation according to claim 1, characterized in that, Injecting the detonating agent into the detonation hole also includes the following steps: The static blasting agent and water are rapidly mixed at a water-cement ratio of 0.3 to form a static blasting agent slurry. The mixing time for each stage is controlled at 1 to 3 minutes. After mixing, the slurry is immediately injected into the static blasting hole in stages using a grouting pump, stopping when the slurry is 1 to 1.5 meters away from the hole opening.

9. The method for preventing coal seam spalling during deep rockburst tunnel excavation according to claim 8, characterized in that, It also includes the following steps: 30-60 minutes after the static detonator slurry is injected, the hydration reaction begins. When the acoustic emission monitoring equipment detects signs of microcrack expansion in the coal body around the static detonator hole, clean water is injected into all free holes at a pressure of 1-3 MPa for 10-20 minutes to replenish the water evaporated during the hydration reaction and ensure that the detonator reacts fully.

10. The method for preventing coal seam spalling during deep rockburst roadway excavation according to claim 8, characterized in that, The sealing of the static explosion hole includes the following steps: filling and sealing the static explosion hole with polyurethane material, with a sealing length of 1~1.5m.

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