Coordinated control method for flushing, pumping and supporting of soft low-permeability outburst coal seam

Abstract

The invention discloses a flushing, pumping and supporting coordination control method for a soft low-permeability outburst coal seam. 0.5, the permeability coefficient lambda < lt >; the gas content gt is 10 <-2 > m < 2 > / (MPa2. D); and the problems of improper coal flushing amount, discordance between extraction and deformation, unreasonable supporting parameters and the like are solved. The method comprises the steps of exploring parameters such as coal seam thickness and gas content; hydraulic punching process parameters are determined, for example, the coal punching amount is 1.5-2.0 t / m, and the drilling distance is 4-6 m; a cavity and an anti-reflection coal seam are formed through layer-penetrating drilling with the diameter ranging from 94 mm to 113 mm and the jet flow pressure ranging from 16 MPa to 20 MPa; coal deformation is monitored, and stabilization time T1 (stress lt; 0.9 time of protolith); when the gas is extracted to reach the standard, time T2 is determined; and T1 and T2 are compared, and tunneling time T3 and supporting parameters are determined. The method significantly improves the extraction efficiency, reduces the roadway deformation rate, significantly reduces the gas overrun rate, achieves safe and efficient treatment, and is suitable for complex geological conditions.

Description

Technical Field

[0001] This invention belongs to the field of coal mine gas control technology, specifically relating to a coordinated control method for flushing, drainage, and support in soft, low-permeability outburst coal seams. Background Technology

[0002] Coal and gas outbursts are one of the main hidden dangers threatening safe production during coal mining. This is especially true for areas with a firmness coefficient (f) less than 0.5 and a permeability coefficient (λ) less than 10. -2 m 2 / (MPa 2 ·d) Gas content exceeding 8m 3 The management of loose, low-permeability coal seams with a pert of / t is extremely difficult. These seams have a loose coal structure and poor permeability, making it difficult for gas to be released effectively. During mining operations, this easily leads to excessive gas concentrations, exceeding of outburst risk prediction indicators (such as S-value and K1 value), and even serious accidents. Statistics show that outbursts related to loose, low-permeability coal seams account for over 30% of coal mine gas accidents in my country, causing not only casualties and economic losses but also severely hindering efficient coal mine production.

[0003] Traditional gas control methods mainly include cross-seam drilling and in-seam drilling for pre-gas extraction. While these measures can partially extract gas, they are limited by the low permeability of the coal seam, resulting in low extraction efficiency, long lead times to achieving the required level, and often causing imbalances in mining operations. To enhance the effect, hydraulic perforation technology has been introduced, which uses high-pressure water jets to create cavities in the coal seam, achieving pressure relief and increased permeability. This technology has been widely applied in mining areas such as Huainan in Anhui, Hancheng in Shaanxi, and Shanxi. For example, in the Huainan mining area, hydraulic perforation can increase the permeability of the coal seam by 2 to 3 times, and the gas extraction rate can be increased to over 50%. However, significant defects still exist in actual implementation: First, the amount of coal flushed is not properly controlled. If it is insufficient, local pressure relief will be inadequate, leaving residual outburst risks; if it is excessive, the coal body will be overly loose, making subsequent coal roadway excavation and support difficult, and easily causing roadway deformation or collapse. Second, the deformation of the coal body after hydraulic perforation is not coordinated with the gas extraction time, and the extraction screen is prone to failure as the coal body deforms, affecting continuous extraction. Third, the support parameters lack scientific basis and do not take into account the deformation stabilization time, resulting in a mismatch between the support strength and actual needs, which increases safety risks.

[0004] While existing technologies attempt to optimize individual aspects, such as adjusting jet pressure or borehole spacing, these are mostly isolated measures that neglect the coupling of multiple factors. For example, some methods focus only on permeability enhancement without quantifying the matching relationship between coal stress recovery and achieving extraction targets, leading to incomplete gas control or excessively high support costs. In low-permeability coal seams, the cavities formed by hydraulic perforation can promote fracture propagation, but without dynamic monitoring, the time it takes for coal stress to recover to less than 0.9 times the original rock level is difficult to accurately determine, thus affecting the timing of tunneling. Therefore, the coal mining industry urgently needs a systematic and coordinated approach that organically integrates perforation, extraction, and support to achieve efficient and safe gas control. Summary of the Invention

[0005] In view of this, the purpose of this invention is to solve the problems in the prior art, such as unreasonable coal flushing volume in hydraulic perforation leading to insufficient pressure relief and permeability enhancement or support difficulties, incoordination between gas extraction and coal deformation, and lack of scientific basis for roadway support parameters. It provides a method for coordinated control of flushing, extraction and support in soft, low-permeability outburst coal seams. By scientifically exploring parameters, optimizing the perforation process, monitoring deformation and extraction time, and dynamically adjusting support strategies, it eliminates outburst hazards, improves gas extraction efficiency, and ensures roadway stability.

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

[0007] A method for coordinated control of flushing, drainage, and support in soft, low-permeability outburst coal seams includes the following steps:

[0008] S1) Drill exploration boreholes in the loose, low-permeability, outburst-prone coal seam to be treated in order to explore and obtain the basic parameters of the coal seam, including thickness, gas content, and firmness coefficient.

[0009] S2) Based on the basic parameters obtained in step S1), determine the process parameters for hydraulic punching operation. The process parameters include at least the amount of coal punched per unit length of the coal hole section and the spacing between punching holes.

[0010] S3) According to the process parameters determined in step S2), hydraulic perforation is carried out in the coal seam to form a pressure relief and permeability enhancement zone;

[0011] S4) Monitor and analyze the evolution of the mechanical state of the coal body in the pressure relief and permeability enhancement area after hydraulic punching operation, and determine the optimal coordinated deformation time T1 of the coal body required for the deformation of the coal body to tend to stabilize.

[0012] S5) Perform gas extraction in the pressure relief and permeability enhancement area, and determine the gas extraction time T2 required to reach the predetermined safety standard;

[0013] S6) Compare the determined T1 and T2 in steps S4) and S5), and jointly determine the reasonable starting time T3 for subsequent drivage of the coal roadway and the corresponding roadway support parameters according to their relative relationship.

[0014] Further, in step S6), the method for determining T3 and the roadway support parameters according to the comparison result of T1 and T2 is as follows:

[0015] When T1 < T2, then T3 ≥ T2, and the first type of support parameter is adopted to meet the requirements of drainage reaching the standard;

[0016] When T1 = T2, then T3 = T1 = T2, and the second type of support parameter is adopted for normal support;

[0017] When T1 > T2, then T1 ≥ T3 ≥ T2, and the third type of support parameter is adopted to enhance the support stability.

[0018] Further, the row spacing of bolts or cable bolts for the first type of support parameter is (8 - 10) m × (8 - 10) m; the row spacing of bolts or cable bolts for the second type of support parameter is (6 - 8) m × (6 - 8) m; the row spacing of bolts or cable bolts for the third type of support parameter is (4 - 6) m × (4 - 6) m.

[0019] Further, in step S4), the basis for determining the optimal coordinated deformation time T1 of the coal body is: when the deformation rate of the coal body in the punching influence area tends to be stable and the coal body stress value recovers to a certain specific threshold lower than the original rock stress, the elapsed time is T1; the specific threshold is 0.9 times the original rock stress.

[0020] Further, the value range of the coal output per unit length of the coal hole section in step S2) is 1.5 - 2.0 t / m.

[0021] Further, the value range of the layout spacing of the punching boreholes in step S2) is 4 - 6 m, and it is adjusted according to the gas content in the coal seam. The higher the gas content, the smaller the layout spacing.

[0022] Further, in step S3), the hydraulic punching operation is to construct cross - layer boreholes in the floor roadway below the coal seam, and use a water jet drilling and expanding integrated device to punch holes in the coal hole section, hollow out the coal body in a certain range around the coal hole section of the cross - layer borehole, and form a cavity. The surrounding coal body moves and deforms towards the cavity, and at the same time drives the fissures to develop, expand and connect to adjacent cavities.

[0023] Further, the diameter of the cross - layer borehole is 94 - 113 mm, the water jet pressure is 16 - 20 MPa, and the coal body in the range of 0.4 - 至0.6 m around the coal hole section of the cross - layer borehole is hollowed out to form a cavity.

[0024] Furthermore, in step S5), gas extraction is carried out by converting the borehole that has completed the punching operation into an extraction borehole, and inserting a flexible screen pipe into the borehole. The screen pipe moves synchronously with the deformation of the coal body to maintain continuous extraction, and the extraction volume of a single hole and the total extraction volume are measured until the total extraction volume reaches the target.

[0025] Furthermore, the process parameters determined in step S2) also include the punching time T per meter, the amount of coal punched per meter of coal hole section is 1.5 to 2.0 t, the punching time T per meter is 10 to 15 min, the punching and drilling spacing is 4 to 6 m, and the spacing is adjusted according to the gas content. The higher the gas content, the smaller the spacing.

[0026] Furthermore, in step S1, the exploration of basic parameters of the coal seam includes: drilling one borehole every 10 to 20 m in the coal seam to measure the thickness of the coal seam, gas content, initial gas release velocity, firmness coefficient, and permeability coefficient.

[0027] The beneficial effects of this invention are as follows:

[0028] Compared with existing technologies, this invention, through systematic process design and innovative coordination and control mechanisms, brings significant technological progress and comprehensive benefits. Its beneficial effects are specifically reflected in the following aspects:

[0029] First, it achieves a paradigm shift from "isolated governance" to "system synergy," constructing a closed-loop control system for safe production. Traditional methods treat drilling, extraction, and support as linear, independent processes, with the connection between processes relying on extensive experience, often resulting in neglecting one aspect for another. This invention, for the first time, places these three stages within a unified systems engineering framework, using geological parameters as input, T1 and T2 as core state variables, and tunneling time and support parameters as outputs, forming a complete closed loop of "exploration-design-construction-monitoring-decision-making." This systematic methodology ensures that decisions at each stage fully consider the impact on subsequent stages, fundamentally resolving the inherent contradictions caused by information fragmentation and differing objectives. It makes the "effectiveness" of gas control and the "long-term effectiveness" of roadway maintenance no longer opposing goals, but rather a unified whole that can be achieved through collaborative optimization, greatly enhancing the scientific nature and internal coherence of the entire coal seam management project.

[0030] Second, the introduction of quantitative time indicators has achieved a major breakthrough from "empirical judgment" to "scientific decision-making", significantly enhancing the predictability and safety of the project. The present invention innovatively proposes and defines two core physical quantities, namely, "the optimal coordinated deformation time T1 of coal body" and "the gas drainage compliance time T2". T2 ensures the red line of "no production without safety", that is, any tunneling operation must be carried out after the gas drainage reaches the standard, which completely eliminates the risk of gas overrun from the time dimension. More importantly, the introduction of T1 quantifies for the first time the fuzzy, dynamic and invisible "black box" process of the mechanical state of the surrounding rock after hydraulic punching through a measurable and determinable time window. It tells the engineering technicians when the stress release and deformation adjustment of the surrounding rock have basically completed and entered a relatively stable "plateau period". Tunneling after T1 means that the stage with the most intense and unstable surrounding rock properties can be avoided, thus greatly reducing the occurrence probability of dynamic disasters such as roadway instability, rib spalling and roof collapse. This data-based dual-time constraint decision-making mechanism replaces the past "arbitrary" empiricism and provides unprecedented scientific basis and safety guarantee for the selection of roadway tunneling timing.

[0031] Third, a differentiated support strategy has been created, realizing the innovation of the support concept from "passive response" to "active pre-control" and bringing significant economic benefits. The most ingenious part of the present invention lies in the decision matrix in step S6, which matches three different support schemes for the roadway according to the different relationships between T1 and T2. When T1 < T2, it indicates that the surrounding rock has already been stable during tunneling. At this time, a relatively loose support spacing can be adopted, which is a "cost optimization" strategy under the premise of ensuring safety, avoiding unnecessary waste of materials and working hours. When T1 > T2, it indicates that although the gas has reached the standard, the surrounding rock is still in an unstable period. At this time, the system will "actively give an early warning" and instruct to adopt encrypted enhanced support. This is a "risk pre-control" strategy. By increasing the upfront investment, it effectively prevents large deformations or instability repairs that may occur later, avoiding costly rework and production interruptions. This refined and dynamic support mode of "taking measures according to different times" and "adopting appropriate measures according to local conditions" makes the support investment accurately match the actual risks, achieving the best balance between safety and economy. It is a subversive improvement to the traditional "one-size-fits-all" support design and has great potential economic value.

[0032] In summary, the present invention is not only a specific technical method, but also a set of advanced engineering management philosophies. It has successfully solved the core pain points in the treatment of soft and low-permeability outburst coal seams through systematic, quantitative and dynamic means, achieving a high degree of unity of safety, efficiency and economy, and has extremely high engineering application value and broad promotion prospects.

[0033] Other advantages, objectives, and features of the invention will be set forth in part in the description which follows, and in part will be apparent to those skilled in the art from the following examination, or may be learned from practice of the invention. The objectives and other advantages of the invention can be realized and obtained through the following description. Attached Figure Description

[0034] To make the objectives, technical solutions, and advantages of the present invention clearer, the preferred embodiments of the present invention will be described in detail below with reference to the accompanying drawings, wherein:

[0035] Figure 1 This is a flowchart illustrating the coordinated control method for flushing, pumping, and supporting soft, low-permeability outburst coal seams in this invention.

[0036] Figure 2 This is a construction schematic diagram of the coordinated control method for flushing, pumping, and supporting soft, low-permeability outburst coal seams in this invention.

[0037] Figure 3 This is the curve showing the deformation and stress change of the coal body after punching in this invention.

[0038] Figure 4 The curves show the changes in single-hole extraction volume versus extraction time.

[0039] Attached diagram labels: 1-coal seam; 2-bottom roadway; 3-through borehole; 4-cavity. Detailed Implementation

[0040] The following specific examples illustrate the implementation of the present invention. Those skilled in the art can easily understand other advantages and effects of the present invention from the content disclosed in this specification. The present invention can also be implemented or applied through other different specific embodiments, and various details in this specification can be modified or changed based on different viewpoints and applications without departing from the spirit of the present invention. It should be noted that the illustrations provided in the following embodiments are only schematic representations of the basic concept of the present invention. Unless otherwise specified, the following embodiments and features can be combined with each other.

[0041] The accompanying drawings are for illustrative purposes only and are schematic diagrams, not actual pictures. They should not be construed as limiting the invention. To better illustrate the embodiments of the invention, some parts in the drawings may be omitted, enlarged, or reduced, and do not represent the actual product dimensions. It is understandable to those skilled in the art that some well-known structures and their descriptions may be omitted in the drawings.

[0042] In the accompanying drawings of the embodiments of the present invention, the same or similar reference numerals correspond to the same or similar components. In the description of the present invention, it should be understood that if terms such as "upper," "lower," "left," "right," "front," and "rear" indicate the orientation or positional relationship based on the orientation or positional relationship shown in the drawings, they are only for the convenience of describing the present invention and simplifying the description, and do not 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, the terms used to describe positional relationships in the drawings are only for illustrative purposes and should not be construed as limiting the present invention. For those skilled in the art, the specific meaning of the above terms can be understood according to the specific circumstances.

[0043] In a preferred embodiment of the present invention, for a soft, low-permeability outburst coal seam in a certain mining area (firmness coefficient f = 0.4, permeability coefficient λ = 5 × 10⁻⁶), -3 m 2 / (MPa 2 ·d) Gas content = 10m 3 / t), this method is applied to coordinate the control of pumping, drawing, and support. The specific steps are as follows:

[0044] S1. Drill exploratory boreholes in the loose, low-permeability, and outburst-prone coal seam to be treated to explore and obtain the basic parameters of the coal seam, including thickness, gas content, and firmness coefficient. Specifically, drill one borehole every 10-20m in coal seam 1, and use drilling equipment to determine the coal seam thickness as 2.5m and the gas content as 10m. 3 / t, initial gas release velocity is 15 mmHg / min, robustness coefficient f = 0.4, air permeability coefficient λ = 5 × 10 -3 m 2 / (MPa 2 •d). These parameters provide the data foundation for subsequent steps, ensuring the relevance of process parameters. This step is completed through on-site drilling and laboratory testing, taking approximately 1–2 days.

[0045] S2. Based on the basic parameters obtained in step S1, determine the process parameters for hydraulic drilling operations. These parameters include at least the coal flushing rate per unit length of the coal seam and the spacing between the drilling holes. In practice, considering the softness of the coal seam and high gas content, the coal flushing rate per unit length of the coal seam is set at 1.5–2.0 t / m (1.8 t / m in this example) to balance the pressure relief effect and support stability; the drilling time T per meter is 10–15 min (12 min in this example) to ensure sufficient but not excessive flushing; the spacing between the drilling holes is 4–6 m (5 m in this example), adjusted according to the gas content—since the gas content is 10 m... 3 / t (relatively high), with a smaller spacing. This parameter was determined and optimized using empirical formulas and numerical simulation software (such as FLAC3D), with the goal of creating an effective pressure relief and anti-reflection zone.

[0046] S3. Following the process parameters determined in step S2, hydraulic perforation is performed in the coal seam to create a pressure-relief and permeability-enhancing zone. Specifically, a cross-layer borehole 3 is constructed in the bottom roadway 2 below coal seam 1, with a borehole diameter of 94–113 mm (100 mm in this example). A water jet drilling and reaming integrated device is used for perforation of the coal seam section, with a jet pressure of 16–20 MPa (18 MPa in this example). After perforation, the coal body within a 0.4–0.6 m (approximately 0.5 m in this example) radius around the coal seam section of the cross-layer borehole 3 is hollowed out to form a cavity 4. The surrounding coal body moves and deforms into the cavity 4, simultaneously driving the development, expansion, and connection of fractures into adjacent cavities, thereby improving the overall permeability of the coal body within the perforation range. Figure 2 As shown, multiple through-layer boreholes 3 are arranged along the bottom roadway 2, forming a three-dimensional pressure relief network. Each group of boreholes takes about 2 to 3 hours to complete this step, with a total coal output of about 4 to 5 tons per borehole.

[0047] S4. Monitor and analyze the evolution of the mechanical state of the coal in the pressure relief and permeability enhancement area after hydraulic drilling, and determine the optimal coordinated deformation time T1 required for the coal deformation to stabilize. In practice, the coal flushing rate per meter (1.8 t / m) and the total coal output per hole are statistically analyzed for all cross-layer boreholes 3. Theoretical calculations (such as elastoplastic mechanical models) and numerical simulation software (such as ANSYS) are used to analyze the deformation and stress variation patterns of the coal. When the coal deformation rate in the drilling influence area stabilizes and the coal stress value recovers to less than 0.9 times the original rock stress (original rock stress is 10 MPa, threshold is 9 MPa in this example), the time elapsed is T1 (T1 = 20 days in this example). Figure 3 As shown in the curve, the deformation initially increases rapidly (0–5 days, with a deformation of 50 mm), then stabilizes, and the stress recovers to 8.5 MPa after initially decreasing to 4 MPa. This step was monitored in real time using stress sensors and displacement gauges, and verified through simulation.

[0048] S5. Perform gas drainage in the pressure relief and permeability enhancement area, and determine the gas drainage compliance time T2 required to reach the predetermined safety standard. Specifically, after completing each set of perforated boreholes, convert the cross-layer borehole 3 into a drainage borehole, and insert a flexible screen pipe (made of flexible plastic composite pipe) throughout the borehole. The screen pipe moves synchronously with the deformation of the coal body to maintain continuous drainage. Measure the drainage volume of each borehole and the total drainage volume, observe the growth trend of the total drainage volume and the overall drainage compliance status (the predetermined standard is that the gas content drops to 6m³). 3 / t or less). When the extraction time reaches T2, the total extraction volume meets the standard (in this example, T2 = 25 days, and the total extraction volume is 500m). 3 ).like Figure 4 As shown in the figure, the curve indicates that the flow rate per orifice is 0.5 m³ / s during the initial stage of extraction.3 / min, and then decays to 0.1 m 3 / min. This step uses a flow meter and a gas concentration monitor to ensure efficient extraction.

[0049] S6. Compare T1 and T2 determined in steps S4 and S5, and collaboratively determine the reasonable starting time T3 for subsequent coal roadway driving and the corresponding roadway support parameters according to their relative relationship. In this example, T1 = 20 days < T2 = 25 days, so T3 ≥ T2 (take T3 = 25 days), and the first support parameter is adopted to meet the requirements of extraction compliance - the row spacing of bolts or cable bolts is (8 - 10) m × (8 - 10) m (take 9 m × 9 m in this example). If T1 = T2, the second support parameter is adopted, with a row spacing of (6 - 8) m × (6 - 8) m for normal support; if T1 > T2, the third support parameter is adopted, with a row spacing of (4 - 6) m × (4 - 6) m to enhance support stability. This step ensures that driving is carried out after the coal body is stable and the gas is up to standard, avoiding risks.

[0050] Through the above embodiments, as Figure 1 shown in the overall process, after this method is applied in this mining area, the gas extraction efficiency is increased by about 30%, the roadway deformation rate is reduced by 12%, and the outburst risk prediction indexes S value and K1 value are not exceeded, achieving coordinated control of punching, extraction, and support. This embodiment is applicable to coal seams with similar geological conditions and can be fine-tuned according to actual parameters.

[0051] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention and not to limit them. Although the present invention has been described in detail with reference to the preferred embodiments, those of ordinary skill in the art should understand that the technical solutions of the present invention can be modified or equivalently replaced without departing from the purpose and scope of the present technical solution, and they should all be covered within the scope of the claims of the present invention.

Claims

1. A method for coordinated control of flushing, pumping and supporting of a soft, low-permeability, outburst coal seam, characterized in that, The method comprises the following steps: S1) constructing an exploration borehole in a soft and low-permeability outburst coal seam to be treated to explore and obtain basic parameters of the coal seam, including thickness, gas content, and firmness coefficient; S2) determining process parameters of hydraulic punching operation according to the basic parameters obtained in step S1), the process parameters at least including coal punching amount per unit length of coal hole section and punching borehole arrangement interval; S3) implementing the hydraulic punching operation in the coal seam according to the process parameters determined in step S2) to form a pressure relief and permeability improvement area; S4) monitoring and analyzing the mechanical state evolution process of the coal body in the pressure relief and permeability improvement area after the hydraulic punching operation to determine the optimal coordinated deformation time T1 of the coal body required for deformation to be stable; S5) performing gas extraction in the pressure relief and permeability improvement area and determining the gas extraction time T2 required for gas extraction to reach a predetermined safety standard; S6) comparing the T1 and T2 determined in steps S4) and S5) and determining the reasonable starting time T3 of subsequent coal roadway excavation and the corresponding roadway support parameters according to the relative relationship between the two.

2. The method of claim 1, wherein, In step S6), the T3 and the roadway support parameters are determined according to the comparison results of the T1 and the T2 as follows: When T1 < T2, T3 ≥ T2, and the first support parameter is adopted to meet the extraction standard requirement; When T1 = T2, T3 = T1 = T2, and the second support parameter is adopted for normal support; When T1 > T2, T1 ≥ T3 ≥ T2, and the third support parameter is adopted to strengthen the support stability.

3. The method of claim 2, wherein, The row and column interval between anchor rods or anchor cables of the first support parameter is (8-10) m x (8-10) m; the row and column interval between anchor rods or anchor cables of the second support parameter is (6-8) m x (6-8) m; and the row and column interval between anchor rods or anchor cables of the third support parameter is (4-6) m x (4-6) m.

4. The method of claim 1, wherein, In step S4), the optimal coordinated deformation time T1 of the coal body is determined as follows: when the coal body deformation rate in the punching influence area tends to be stable and the coal body stress value recovers to a certain specific threshold value lower than the original rock stress, the time experienced is T1; the specific threshold value is 0.9 times the original rock stress.

5. The method of claim 1, wherein, In step S2), the coal punching amount per unit length of coal hole section is in the range of 1.5-2.0 t / m.

6. The method of claim 1, wherein, In step S2), the arrangement interval of the punching borehole is in the range of 4-6 m and is adjusted according to the coal seam gas content; the higher the gas content, the smaller the arrangement interval.

7. The method of claim 1, wherein, In step S3), the hydraulic punching operation is performed by constructing a through-hole borehole in the target coal seam in a floor roadway below the coal seam and using a water jet drilling and expanding integrated device to punch the coal hole section, thereby hollowing out the coal body within a certain range around the coal hole section of the through-hole borehole and forming a cavity, and the surrounding coal body moves and deforms into the cavity, and the cracks develop, expand and connect to the adjacent cavities.

8. The method of claim 7, wherein, The through-hole borehole has a diameter of 94-113 mm, and the water jet pressure is 16-20 MPa, which can hollow out the coal body within a range of 0.4-0.6 m around the coal hole section of the through-hole borehole and form a cavity.

9. The method of claim 1, wherein, In step S5), gas extraction is performed by transforming the completed drilling into an extraction drilling, and inserting a ductile screen pipe into the drilling, which moves synchronously with the deformation of the coal body, maintains continuous extraction, measures the single-hole and total extraction amount, and stops until the total extraction amount reaches the standard.

10. The method of claim 1, wherein, The process parameters determined in step S2) further include the drilling time T per meter, the coal drilling amount per meter is 1.5-2.0 t, the drilling time T per meter is 10-15 min, the drilling interval is 4-6 m, and the interval is adjusted according to the gas content, the higher the gas content, the smaller the interval.

11. The method of claim 1, wherein, In step S1, the basic parameters of the coal seam are explored, including: constructing a drilling every 10-20 m in the coal seam to measure the thickness, gas content, gas diffusion initial velocity, firmness coefficient, and gas permeability coefficient of the coal seam.

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