A research on destress mechanism of underlying roadway for gas outburst prevention in deep, low-permeability coal seam

Abstract

Gas outbursts pose a significant threat in the mining of deep low-permeability coal seams. As mining depth increases, there is a rise in coal and gas dynamic disasters, rendering the gas outburst prevention technology from shallow mining areas unsuitable for deep coal seam exploitation. Addressing the challenges posed by deep, highly gassy coal seams, this study introduces a gas outburst prevention technology integrating destressing in the close distance underlying roadway with gas drainage. The article investigates the relevant technical parameters and assesses the feasibility of this approach. First, it establishes destressing model for the underlying roadway to derive an analytical solution for stress distribution in the coal seam. Second, appropriate technical parameters are designed considering the seepage characteristics of gas in the coal seam. Finally, the gas drainage technical scheme is implemented at Qujiang coal mine to verify its on-site effectiveness. Field test results indicate an increase in gas permeability within the coal seam following destressing in the underlying roadway, accompanied by significant improvements in the efficiency of gas drainage and the advancement speed of the coal roadway. Specifically, the driving speed of the coal roadway has escalated from 40 m/month to 150 m/month. These outcomes demonstrate the potential of this technology as a promising approach for preventing gas outbursts in the challenging context of deep, low-permeability coal seams, facilitating swift coal roadway excavation.

INTRODUCTION

Coal resources play a pivotal role in global industrial energy development.¹ As coal mining extends deeper, the gas environment exhibits characteristics such as elevated ground stress, high gas content and pressure, and low permeability, heightening the risk of gas outbursts.^2–5 During coal seam mining, the stress equilibrium in the rock mass is disrupted, potentially leading to sudden outflow, ignition, or explosion of methane, outburst of rocks, and methane.^6–8 Geological factors, including continuous fault disturbances and coal properties such as strength, sorption capacity, desorption kinetics, and initial gas pressure, are believed by many scholars to be the primary causes of coal and gas outbursts.^9–11 Gas accidents pose significant safety hazards in coal mining, and the casualties resulting from gas outbursts during deep coal seam mining worldwide need to be taken seriously.^1,12 However, gas remains integral to the global energy supply as a clean energy source.¹³ Gas predrainage and destressing are common preventive measures against gas outbursts. Gas pre-drainage not only prevents energy waste but also ensures safety in coal seam mining, thereby facilitating technical innovation in the simultaneous mining of coal and gas.^14,15

Permeability of coal seams is a crucial parameter in determining the effectiveness of gas drainage from coal seams.¹⁶ In engineering, destressing technology is employed to alleviate stress in coal seams. This process enhances the permeability of gas-containing coal seams, consequently mitigating the challenges associated with gas drainage.¹⁷ Protective layer mining is typically integrated with destressing gas drainage and pre-drainage of coal seam gas through large-scale, densely spaced drilling.^18–20 The coal roadway strip within the protected layer is commonly achieved through the simultaneous application of destressing pumping and mining activities. Hydraulic techniques aimed at enhancing the gas permeability of coal seams primarily consist of hydraulic punching,^21–23 hydraulic slotting,^24,25 and hydraulic fracturing.^26–28 A fracturing pump assembly is employed to inject high-pressure water into a sealed hole. This process induces fracturing fractures within the original coal body, establishing a flow channel for gas movement. Blasting aimed at enhancing the gas permeability of the coal seam involves utilizing high temperature and high pressure generated by explosive explosions to impact and fracture the coal body, thereby creating an extensive network of cracks.^29–32 At the same time, existing cracks in the coal body are cut through, forming a fracture network that serves as a channel for gas migration, thereby enhancing the gas drainage efficiency of drilling.

Nevertheless, for individual coal seams with high deep gas content and low permeability, adopting the protective layer mining method to achieve destressing and enhanced gas permeability is considered unfeasible.²⁵ In instances where coal seams exhibit weak lithology, the hydraulic fracturing method proves ineffective in achieving the expected outcomes in terms of gas drainage.³³ In cases where a single coal seam is characterized by high gas content and soft composition, engineering often employs the pre-cracking blasting method to enhance the permeability of the coal seam in the section of the coal roadway slated for excavation. Blasting failures surrounding the borehole serve not only to diminish stress concentration in the vicinity but also to augment coal seam permeability through coal body failure, thereby enhancing the gas drainage rate.^30,31 However, this method exhibits several limitations: challenges in drilling deep into soft coal seams, inadequate hole formation effectiveness, difficulties in charging, and a high energy absorption effect in soft coal seams. Moreover, cracks easily compact under the influence of high ground stress, resulting in suboptimal outcomes for destressing through blasting and an increase in seepage. In situations where the upper and lower rock layers exhibit significant fracturing or are located in close proximity to geological structural zones, the propagation of blasting stress waves intensifies the fragmentation of rock layers. This presents challenges in ensuring optimal roof support quality during mining operations and introduces potential safety hazards.³⁴

The paper proposes a gas outburst prevention technology for deep coal seam mining, which integrates destressing of underlying roadways with layered drilling for gas drainage. By analyzing the destressing law of close-distance underlying roadway and combining it with the relevant characteristics of coal seam gas permeability, the layout parameters for underlying roadway are determined, and field industrial experiments were conducted. The primary objective of this study is to introduce a new approach to prevent and control gas outbursts in deep coal mining.

GEOLOGICAL BACKGROUND AND TECHNOLOGY SYNOPSIS Geological background

The Qujiang Coal Mine, situated in Fengcheng City, Jiangxi Province, serves as the primary mine in the Fengcheng Mining area, as depicted in Figure 1. The B₄ coal seam, with an average thickness of 3.3 m, is the primary focus of mining operations. The roof of the coal seam is predominantly composed of fine sandstone and mudstone, exhibiting effective sealing. Figure 2 illustrates the local column shape of the coal measure strata. The buried depth of the B₄ coal seam reaches −850 m, featuring high gas content, with a maximum gas pressure of 9.2 MPa and a gas content ranging from 13.5 to 25.3 m³/t. Due to its low permeability and challenges associated with gas drainage efficiency, extracting gas from this coal seam presents significant obstacles for effective drainage processes. Over the years, Fengcheng Mining area has witnessed 186 gas outburst accidents, as depicted in Figure 1. Since the commencement of operations at Qujiang Coal Mine, there have been eight outburst incidents, comprising five significant gas outbursts, and three moderate gas outbursts. The maximum coal amount involved in an outburst is 860 t, with a maximum gas emission of 36,000 m³. The gas outburst intensity is pronounced and escalates with increasing mining depth, particularly during coal roadway excavation.

Figure 1. Location of Qujiang coal mine.

Figure 2. Rock column diagram.

Technology synopsis

Excavating a close-range underlying roadway to release the coal seam pressure, combined with rock drilling and gas drainage technology, is an effective approach to prevent gas outburst. The specific principles are illustrated in Figure 3. Initially, the gas permeability of the coal seam is enhanced by excavating the underlying roadway for local destressing. Subsequently, gas drainage is executed through cross-layer drilling to lower the gas pressure and content in the excavation area to meet safety permit requirements. Consequently, eliminating the hazards of gas outbursts during coal mining is beneficial for the safe and efficient excavation of coal seam roadway. This technology carries significant economic and technical advantages in addressing gas outbursts in deep coal seam areas.

Figure 3. Engineering principle.

THEORETICAL FUNDAMENTALS

This study involves the establishment of a destressing model for the underlying roadway, wherein the analysis delves into the stress variation in the coal seam concerning distinct underlying roadway radii and layout distances. By considering the gas seepage characteristics of the B₄ coal seam, the investigation scrutinizes the variation in gas permeability within the coal seam corresponding to different positions of the underlying roadway beneath it. Ultimately, this research conclusively determines optimal design parameters such as radius and layout distance for the underlying roadway.

Mechanical analysis model of roadway

In analyzing the stress distribution of roadway surrounding rock, it is common practice to simplify the roadway section into a circular shape using the method of fitting an external circle.³⁵ According to the stress-strain curve of rocks,³⁶ the stress distribution in the surrounding rock after roadway excavation can be classified into three states: elastic zone, plastic failure zone, and fracture zone. The mechanical model for the surrounding rock of deep roadways, as depicted in Figure 4, can be established.

Figure 4. Mechanical analysis model of roadway.

Basic theory and equations

To facilitate the theoretical analysis, the roadway cross-section can be approximated as a circular cross-section, and the deep surrounding rock is assumed to be uniform, isotropic, and in a plane strain state. The roadway has a radius denoted as R₀, with an initial ground stress of p₀ and a support stress of p₁. The surrounding rock comprises three zones: the rupture zone, plastic zone, and elastic zone, illustrated in Figure 4. Subscripts “c”, “p,” and “e” represent variables in each respective zone. Using polar coordinates, radial and circumferential stress and strain are denoted by σ_r, σ_θ, and ε_r, ε_θ, while displacement is denoted by u.

The equilibrium differential equation governing stress in each zone is (excluding consideration of volume force): [Image Omitted. See PDF]

The geometric equation is formulated as follows: [Image Omitted. See PDF]

The excavation and unloading of the roadway lead to a support pressure lower than the ground stress, triggering stress redistribution and causing the surrounding rock to yield. When the surrounding rock undergoes plastic yield and fracture, its stress aligns with the M-C strength theory criterion, expressed as: [Image Omitted. See PDF]where, $<math mathsize="9pt"><mrow><mrow><msub><mi>\unicode{x003B7}</mi><mi>i</mi></msub><mo>\unicode{x0003D}</mo><mo>(</mo><mn>1</mn><mo>\unicode{x0002B}</mo><mspace width="0.25em"/><mi>sin</mi><mspace width="0.25em"/><msub><mi>\unicode{x003C6}</mi><mi>i</mi></msub><mo>)</mo><mo>/</mo><mo>(</mo><mn>1</mn><mo>\unicode{x02010}</mo><mspace width="0.25em"/><mi>sin</mi><mspace width="0.25em"/><msub><mi>\unicode{x003C6}</mi><mi>i</mi></msub><mo>)</mo></mrow></mrow></math>$, $<math mathsize="8.8pt"><mrow><mrow><msub><mi>\unicode{x003BE}</mi><mi>i</mi></msub><msub><mrow><mo>\unicode{x0003D}</mo><mtext>2c</mtext></mrow><mi>i</mi></msub><mspace width="0.25em"/><mi>cos</mi><mspace width="0.25em"/><msub><mi>\unicode{x003C6}</mi><mi>i</mi></msub><mo>/</mo><mo>(</mo><mn>1</mn><mo>\unicode{x02212}</mo><mspace width="0.25em"/><mi>sin</mi><mspace width="0.25em"/><msub><mi>\unicode{x003C6}</mi><mi>i</mi></msub><mo>)</mo><mo>.</mo></mrow></mrow></math>$

The boundary conditions are: [Image Omitted. See PDF]

Analytical solutions Elastic region (R_p ≤ r ﹤ ∞)

By utilizing Equations (1), (2), and (3), along with the boundary conditions specified in Equation (4), the following equation can be obtained: [Image Omitted. See PDF]where G represents the shear modulus of rock, and $<math ><mrow><mrow><msub><mi>\unicode{x003C3}</mi><msub><mi>R</mi><mi>p</mi></msub></msub></mrow></mrow></math>$ denotes the contact stress at the elastic-plastic junction. When r = R_p, the stress of the surrounding rock satisfies the critical yield criterion of M-C strength theory. Substituting $<math ><mrow><mrow><msubsup><mi>\unicode{x003C3}</mi><mi>r</mi><mi>e</mi></msubsup></mrow></mrow></math>$ and $<math ><mrow><mrow><msubsup><mi>\unicode{x003C3}</mi><mi>\unicode{x003B8}</mi><mi>e</mi></msubsup></mrow></mrow></math>$ from Equation (5) into Equation (3) yields the equation for $<math ><mrow><mrow><msub><mi>\unicode{x003C3}</mi><msub><mi>R</mi><mi>p</mi></msub></msub></mrow></mrow></math>$: [Image Omitted. See PDF]

Plastic zone (R_c ≤ r ≤ R_p)

By combining Equations (1) and (3) and the boundary conditions in Equation (4), the stress in the plastic zone can be obtained: [Image Omitted. See PDF]where $<math ><mrow><mrow><msub><mi>Q</mi><mi>p</mi></msub><mo>\unicode{x0003D}</mo><msub><mi>\unicode{x003BE}</mi><mi>p</mi></msub><mo>/</mo><mo stretchy="false">(</mo><msub><mi>\unicode{x003B7}</mi><mi>p</mi></msub><mo>\unicode{x02212}</mo><mn>1</mn><mo stretchy="false">)</mo></mrow></mrow></math>$.

After plastic yield, the surrounding rock exhibits noticeable dilatancy characteristics, with a nonzero plastic volume strain, thereby adhering to the linear non-correlated flow rule³³: [Image Omitted. See PDF] where K_p = (1 + sinψ_p)/(1 − sinψ_p), ψ_p is the dilation angle of plastic region, under noncorrelated flow rule, the dilatancy angle should be ψ_{p =} φ_p/2.³⁷

By connecting Equations (2), (3), (4), and (8), the deformation equation can be obtained: [Image Omitted. See PDF] where $<math ><mrow><mrow><msub><mi>N</mi><mn>1</mn></msub><mo>\unicode{x0003D}</mo><mo>(</mo><msub><mi>\unicode{x003C3}</mi><msub><mi>R</mi><mi>p</mi></msub></msub><mo>\unicode{x02212}</mo><msub><mi>P</mi><mn>0</mn></msub><mo>)</mo><mo>(</mo><mn>1</mn><mo>\unicode{x02212}</mo><msub><mi>K</mi><mi>p</mi></msub><mo>)</mo><mo>/</mo><mo>[</mo><mn>2</mn><mi>G</mi><mo>(</mo><mn>1</mn><mo>\unicode{x0002B}</mo><msub><mi>K</mi><mi>p</mi></msub><mo>)</mo><mo>]</mo></mrow></mrow></math>$, $<math ><mrow><mrow><msub><mi>N</mi><mn>2</mn></msub><mo>\unicode{x0003D}</mo><mo>(</mo><msub><mi>\unicode{x003C3}</mi><msub><mi>R</mi><mi>p</mi></msub></msub><mo>\unicode{x02212}</mo><msub><mi>P</mi><mn>0</mn></msub><mo>)</mo><mo>/</mo><mo>[</mo><mi>G</mi><mo>(</mo><mn>1</mn><mo>\unicode{x0002B}</mo><msub><mi>K</mi><mi>p</mi></msub><mo>)</mo><mo>]</mo></mrow></mrow></math>$.

Fracture zone (R₀ ≤ r ≤ R_c)

Combined with the boundary conditions in Equation (4), the stress in the fracture zone can be obtained as follows: [Image Omitted. See PDF] where $<math ><mrow><mrow><msub><mi>Q</mi><mi>c</mi></msub><mo>\unicode{x0003D}</mo><msub><mi>\unicode{x003BE}</mi><mi>c</mi></msub><mo>/</mo><mo stretchy="false">(</mo><msub><mi>\unicode{x003B7}</mi><mi>c</mi></msub><mo>\unicode{x02212}</mo><mn>1</mn><mo stretchy="false">)</mo></mrow></mrow></math>$.

Assuming that the total strain in the fracture zone consists solely of plastic strain, the relationship between circumferential and radial strain in this zone is as follows: [Image Omitted. See PDF] where K_c = (1 + sinψ_c)/(1 − sinψ_c), ψ_c is the dilatancy angle of the fracture zone, the value of the parameter is ψ_c = φ_c/2.³⁷

The deformation equation can also be obtained as: [Image Omitted. See PDF]

Equation solving

By combining the boundary conditions in Equation (9) with Equations (5) and (10), the following relationship can be derived: [Image Omitted. See PDF]

Solving Equation (13) requires the inclusion of an additional equation³⁶: [Image Omitted. See PDF]

By connecting the lines Equations (9), (12), (13), and (14), the equation can be obtained: [Image Omitted. See PDF] [Image Omitted. See PDF] where $<math ><mrow><mrow><mi>\unicode{x003BB}</mi><mo>\unicode{x0003D}</mo><msup><mrow><mo stretchy="false">{</mo><mrow><msub><mi>N</mi><mn>1</mn></msub><mo stretchy="false">(</mo><mn>1</mn><mo>\unicode{x0002B}</mo><msub><mi>K</mi><mi>c</mi></msub><mo stretchy="false">)</mo><mo>/</mo><mrow><mo stretchy="false">[</mo><mrow><msub><mi>N</mi><mn>2</mn></msub><mo stretchy="false">(</mo><msub><mi>K</mi><mi>c</mi></msub><mo>\unicode{x02212}</mo><msub><mi>K</mi><mi>p</mi></msub><mo stretchy="false">)</mo></mrow><mo stretchy="false">]</mo></mrow></mrow><mo stretchy="false">}</mo></mrow><mrow><mn>1</mn><mo>/</mo><mrow><mo stretchy="false">(</mo><mrow><msub><mi>K</mi><mi>p</mi></msub><mo>\unicode{x0002B}</mo><mn>1</mn></mrow><mo stretchy="false">)</mo></mrow></mrow></msup></mrow></mrow></math>$.

Solving the destressing model of coal seam

Employing the coordinate transformation within elastic mechanics, illustrated in Figure 5, the stress change formula in the coal seam under the Cartesian coordinate system is obtained. In the case of axisymmetric problems, the coordinate transformation formula can be simplified as: [Image Omitted. See PDF]

Figure 5. Coal seam unloading mechanics model.

Where α is the angle between a point in the coal seam and the direction of the y axis., σ_y is the vertical stress in the Cartesian coordinate system.

The coal seam should be positioned within the elastic zone of the underlying roadway relief, aligning with the code requirements and project safety considerations during the design of underlying roadway layout spacing. Consequently, by substituting Equation (5) into Equation (17), the stress distribution in the coal seam can be derived as follows: [Image Omitted. See PDF]where x and y are the coordinates of any point in the coal seam.

Project case analysis and discussion

The B₄ coal seam in Qujiang coal mine reaches a depth of approximately 970 m with a coal seam thickness of 3.3 m, and it is predominantly surrounded by fine and coarse sandstone. The primary rock stress p₀ = 24.6 MPa, the elastic modulus of the rock is E = 15 GPa, the Poisson ratio of the rock is μ = 0.25, the initial friction angle φ_p = 36°, and the initial cohesion c_p = 3.2 MPa. Residual internal friction angle φ_c = 12°, residual cohesion c_c = 0.8 MPa, the support resistance p₁ = 0.1 MPa.

Stress distribution of roadway surrounding rock

The equations derived theoretically in Section 3.3 of the paper should be employed for calculation and analysis. Calculations based on the provided parameters reveal that the fracture zone radius R_c of the 213 underlying roadway is 1.9R₀, and the plastic zone radius R_p is 2.9R₀. Figure 6 illustrates the stress-strain curve and surrounding rock parameters of the roadway. To ensure gas drainage tightness and prevent gas leakage through surrounding rock cracks, the underlying roadway should be positioned in zone III. This involves siting the coal seam in the elastic zone of destressing for the underlying roadway, as depicted in Figure 6, ensuring normal and safe gas drainage.

Figure 6. Stress–strain curve of roadway surrounding rock.

Underlying roadway radius analysis

Two pivotal parameters affecting the efficacy of engineering implementation in underlying roadway destressing for coal seams are the roadway radius and layout distance. The prescribed minimum normal distance from the coal seam is 7 m when addressing local outburst.³⁸ To scrutinize the influence of roadway radius on destressing, a layout distance of h = 8 m is employed.

Figure 7 demonstrates a gradual decrease in destressing effect from the bottom to the top of the coal seam, accompanied by an increase in destressing width. When the roadway radius R₀ is 1.5 m, there is a minimum stress of 20.3 MPa beneath the coal seam and 22.2 MPa above it, resulting in an indistinct destressing effect. With an increased roadway radius, the destressing effect is enhanced, while the destressing width remains relatively constant at approximately 10 m, as illustrated in Figure 8a,b. In high ground stress environments, a larger roadway size corresponds to increased support challenges and higher associated costs.

Figure 7. Influence of different radius of roadway on stress distribution of coal seam: (A) R0 = 1.5 m; (B) R0 = 2.0 m; (C) R0 = 2.5 m; (D) R0 = 3.0 m.

Figure 8. Coal seam stress distribution: (A) stress distribution on the lower surface of coal seam; (B) stress distribution on the upper surface of coal seam.

Consequently, a larger underlying roadway size is not universally advantageous; instead, design considerations should align with actual engineering requirements. Balancing the coal seam destressing effect and deep roadway support costs, a roadway radius between 2 and 2.5 m is deemed more appropriate. Drawing from past rock roadway support experiences in Qujiang coal mine, the underlying roadway is designed with a width of 4.8 m, a height of 2.9 m, and an equivalent radius of 2.2 m.

Underlying roadway layout distance analysis

The spacing of the underlying roadway significantly influences the increase in permeability during coal seam destressing. The stress distribution in coal seams under various underlying roadway layout spacings is illustrated in Figures 9 and 10.

Figure 9. Influence of different underlying roadway distance on stress distribution of coal seam: (A) h = 8 m; (B) h = 10 m; (C) h = 12 m; (D) h = 14 m.

Figure 10. Coal seam stress distribution: (A) stress distribution on the lower surface of coal seam; (B) stress distribution on the upper surface of coal seam.

The destressing effect in the coal seam gradually diminishes from bottom to top, as illustrated in Figure 9, accompanied by an increase in the width of destressing. There exists a negative correlation between the destressing amplitude in the coal seam and the layout distance, while a positive correlation is observed with the destressing width. For example, at h = 8 m, the maximum destressing range within the coal seam is 32.5%, with an effective destressing range of 20.4–27 m; at h = 14 m, the maximum destressing range in the coal seam is 13%, and the effective destressing range is 32.4–39 m, as shown in Figure 10. In engineering applications, it is essential to comprehensively consider the impact of coal seam destressing on permeability increase and the width of effective destressing.

Permeability model

Following the excavation of the underlying roadway, there is a noticeable increase in the permeability of the coal seam relief zone. The correlation between gas permeability and stress during coal seam unloading is typically articulated by the following formula^39,40: [Image Omitted. See PDF] where K is the permeability of gas, K₀ is the initial permeability of gas, a is a constant, and Δσ is the stress change value.

Following the sampling of the B₄ coal seam, a destressing permeability test was conducted on coal and gas under triaxial stress. The resulting gas permeability equation for coal samples from the B₄ coal seam during destressing is expressed as follows⁴¹: [Image Omitted. See PDF]

Substituting Equation (18) into formula (20) yields the equation for coal seam permeability as follows: [Image Omitted. See PDF]

The law of permeability increase in coal seam unloading

To determine the optimal layout distance for the underlying roadway, an analysis was conducted to assess the permeability variation of the coal seam at various distances from the roadway.

To directly evaluate the improvement in coal seam permeability, Figure 11 presents a cloud chart value indicating the magnitude of permeability enhancement. The cloud chart in Figure 11 represents the ratio of permeability after coal seam destressing to initial permeability. Notably depicted in the figure is the region where values surpass 1, representing the effective range of enhanced permeability. At a underlying roadway layout distance of h = 8 m, the maximum gas permeability in the coal seam increases to 1.83 times the initial state, with an effective destressing range of 20.4–27 m. Similarly, at h = 14 m, the permeability rises to 1.27 times the initial state, and the effective destressing range expands to 32.4–39 m. It is evident that as the roadway layout distance increases, the enhancement effect on the gas permeability of the coal seam gradually diminishes, while the effective width increases, as illustrated in Figure 12. Therefore, in designing the layout distance of the underlying roadway, it is necessary to comprehensively consider not only increasing coal seam permeability but also ensuring effective width of gas permeability expands.

Figure 11. Influence of different underlying roadway distance on gas permeability of coal seam: (A) h = 8 m; (B) h = 10 m; (C) h = 12 m; (D) h = 14 m.

Figure 12. Variation law of gas permeability in coal seam: (A) permeability of lower surface of coal seam; (B) permeability of upper surface of coal seam.

Considering the magnitude of permeability variation and the effective width of gas permeability expansion, a distance h = 8–12 m from the coal seam is considered more appropriate. In other words, for optimal placement of the underlying roadway, it should be located between 8 and 12 m below the coal seam.

FIELD APPLICATION Layout distance of underlying roadway

To tackle the challenge of gas outbursts in the B₄ coal seam, the proposed method involves depressurizing the coal seam and enhancing permeability through the excavation of a closely positioned underlying roadway. An evolution equation describing gas permeability during coal seam destressing is established, and analyze the impact of different underlying roadway layout spacings on gas permeability. Adhering to the specific engineering requirements of Qujiang coal mine, the underlying roadway is designed with dimensions of 4.2 m width and 2.9 m height, positioned 10 m below the coal seam, as illustrated in Figure 13.

Figure 13. Design of underlying roadway drainage gas.

Under these conditions, the effective destressing range in the horizontal direction spans 24.4 m, with the maximum permeability increase reaching 1.53 times the initial coal seam permeability. Gas drainage in the region where gas permeability is enhanced due to coal seam destressing satisfies the requirements for coal seam roadway excavation. The underlying roadway layout with a 10 m spacing conforms to the gas control code requirements and ensures that the coal seam remains within the elastic zone of stress distribution in the surrounding rock. This effectively prevents gas leakage through cracks in the surrounding rock, thereby ensuring secure project execution.

Parameter design of gas drainage borehole

To ascertain the ideal spacing for gas drainage boreholes, the gas reserve method is applied to compute the effective drainage radius of the initial coal seam.⁴² Gas drainage utilizes a 13 kPa negative pressure with a borehole diameter of Φ96 mm. The gas flow and the effective radius of borehole drainage are illustrated in Figure 14. The curve illustrates that the effective influence radius of drainage for 15d, 30d, 60d, and 90d is 1.12, 1.56, 1.83, and 2.12 m, respectively. The spacing between gas drilling holes is determined based on the effective influence radius of drainage⁴³: [Image Omitted. See PDF]where $<math ><mrow><mrow><mi>R</mi><mo accent="false">\unicode{x02032}</mo></mrow></mrow></math>$ is the spacing between the layout of gas drainage, $<math ><mrow><mrow><mi>r</mi><mo accent="false">\unicode{x02032}</mo></mrow></mrow></math>$ is the effective drainage radius of gas drainage.

Figure 14. Gas flow and effective radius of coal seam borehole drainage.

By analyzing the destressing range and effective gas drainage radius, nine gas drilling holes are designed with a diameter of Φ96 mm and a spacing of 3 m. The gas drainage range is 24 m, meeting engineering requirements and accomplishing the goal of preventing and controlling gas outburst in the coal seam. The specific layout of gas drainage boreholes is depicted in Figure 14. The ventilation system utilizes a combination of central parallel and diagonal mixed drainage ventilation. The main mine shaft and auxiliary mine shaft are used for air intake, while the central and the eastern mine shafts are designated for air return. The ventilator installed in the central mine shaft is the BDK-8-NO32 countercyclone, rated at 630 kW with a negative pressure of 3.2 kPa. Similarly, the eastern mine shaft is equipped with a type FBCDZ-8-NO26 countercyclone fan, rated at 355 kW with a negative pressure of 1.3 kPa. The total air intake of the mine is 11,643 m³/min, with a total air return of 11,983 m³/min. Specifically, the central shaft returns 9892 m³/min, while the eastern shaft returns 2091 m³/min.

To achieve the drilling layout distance, the ZYL-7000D drilling rig is employed, adjusting azimuth and inclination angles as necessary. A Φ96 mm drill bit and Φ73 mm drill pipe are utilized to bore holes into the coal seam. Subsequently, grouting and sealing pipes are inserted into the borehole, followed by sealing its termination with polyurethane. Upon expansion and solidification, the prepared cement slurry is injected through the grouting pipe, completing the sealing process. It is crucial to maintain an appropriate grouting speed to ensure the thorough infiltration of slurry into surrounding cracks, thereby guaranteeing the reliability of the sealed hole. The designed sealing length is approximately 8 m, ensuring the airtightness of the borehole and enhancing gas drainage efficiency.

Gas drainage radius after coal seam destressing

The 213 underlying roadway of Qujiang coal mine is situated 10 m below the coal seam. A rock formation detection recorder, model YTJ20, is employed to observe the internal damage of the roadway's surrounding rock, as illustrated in Figure 15. The drilling view reveals that 1 m above roadway is the theoretical fracture zone, showing evident rock breakage and fully developed cracks. At 3 m above the roadway lies the theoretically calculated plastic zone, with developing cracks and better integrity than at 1 m. The region 6 m above the roadway represents the theoretically calculated elastic zone, where cracks are not developed, and the surrounding rock exhibits good integrity. This observation validates the reliability of the theoretical analysis.

Figure 15. Roof borehole observation and gas drainage borehole layout.

To analyze the change in gas drainage radius after destressing, the gas reserve method was used to determine the effective drainage radius of the coal seam after destressing, as shown in Figure 15. Investigating the variation in gas drainage at horizontal distances of 0, 3, 6, 9, and 12 m elucidates the average gas drainage flow and its corresponding effective radius as shown in Figure 16. The curve illustrates that the position directly above the underlying roadway yields the best gas drainage effect, with a maximum coal seam gas drainage radius of 2.67 m and a maximum daily average gas drainage volume of 32.7 m³, doubling the initial coal seam drainage volume. This demonstrates a substantial increase in gas recovery following destressing through the underlying roadway, thereby confirming the feasibility of this technology. With increasing distance, the gas extracted gradually decreases. At x = 12 m, the amount of gas extracted from the coal seam exceeds that of the initial coal seam, and the permeability enhancement effect of the extracted gas is moderate, confirming the reliability of the theoretical analysis results.

Figure 16. Gas drainage flow and effective drainage radius.

Gas outburst prevention and control effect

Following the destressing of the B₄ coal seam, Figure 17 illustrates the change curve of daily gas drainage and gas concentration. The figure reveals that the maximum concentration of gas extracted by the borehole through the underlying roadway is 14%. The daily gas drainage can reach 550–1538 m³/d, demonstrating a remarkable drainage effect. The driving speed of the coal roadway has significantly increased, from 40 m/month to 150 m/month. After 40 days of gas drainage, the gas content of the coal seam is 3.25–4.93 m³/t, meeting the requirements of 8 m³/t in the outburst prevention code. This achievement realizes the goal of preventing gas outburst in coal seams, providing assurance for the safety of coal roadway excavation.

Figure 17. Gas drainage quantity and drainage concentration.

LIMITATION OF THIS STUDY

The research results indicate that the combination of near-distance underlying roadway unloading and cross-layer drilling gas drainage is effective in regional outburst prevention. This technique can significantly improve gas drainage efficiency and significantly reduce the risk of gas outburst accidents. However, this technique still has certain limitations in practice. In this study, the roadway was located in sandstone with good lithological conditions, making maintenance easier. However, for roadways located in poor lithological layers, it is necessary to analyze the roadway support design and stability when using this technique. In addition, the mining technology of high-gas coal seams affects the deformation law of rock layers,^1,44 thereby influencing the control of residual gas in goaf. Further research is still needed on measures for controlling gas in goaf.

CONCLUSIONS

This article proposes a technique that combines close distance underlying roadway destressing with gas drainage to prevent gas outburst accidents in deep coal seams with high gas content and low permeability characteristics. A mechanical model for underlying roadway destressing is established, and the impact of various parameters on the effectiveness of coal seam destressing is analyzed. Utilizing the gas permeation characteristics of the B₄ coal seam in Qujiang Coal Mine, it determines reasonable technical parameters and verifies their effectiveness through field application. The main conclusions are as follows:

• (1)

  The effective increase range of coal seam gas permeability obtained through theoretical analysis aligns well with the field measurement data. This indicates that the established mechanical model is accurate and it is reasonable to analyze the coal seam destressing effect by simplifying the cross-section of roadway.

• (2)

  As the distance between the underlying roadway and coal seam increases, the enhancement of gas permeability in the coal seam diminishes, but the influence range expands. Therefore, when designing engineering parameters, it is crucial to consider both the magnitude of gas permeability enhancement and the extent of its influence within the coal seam.

• (3)

  Field data demonstrate that coal seam destressing doubles the daily average gas drainage per hole and significantly boosts gas permeability. This technology creates favorable conditions for gas desorption and drainage, providing new technical support for preventing gas outbursts in deep, high-gas, low-permeability coal seam mining.

ACKNOWLEDGMENTS

This work was supported by the National Natural Science Foundation of China (Grant no. 51674247).

CONFLICT OF INTEREST STATEMENT

The authors declare no conflicts of interest.