Research and engineering practice of coupled unloading and supporting technology for rock burst prevention in hard coal seams of the Ordos region

Hard coal seams in the Ordos region are key geological factors contributing to rock bursts. To overcome the limitations of conventional drilling pressure-relief techniques—such as insufficient unloading efficiency, reliance on high-density multi-round drilling, and support failure—this study establishes a non-isobaric stress field model and analyzes the influence of coal seam strength and drilling diameter on the radius of the drilling-induced plastic zone. Based on this analysis, a coupled “shallow support and deep pressure relief” unloading-support technology was proposed. A 70-m on-site comparative industrial test was conducted using coal-powder monitoring, coal-cannon monitoring, stress monitoring, and surrounding-rock deformation monitoring to evaluate pressure-relief and support performance. The results show that the plastic zone radius is mainly controlled by coal strength and borehole diameter, with diminishing benefits when enlarging the diameter beyond a threshold. The enhanced pressure-relief zone produced 3.1 times more coal powder than traditional drilling, and coal-cannon events concentrated in the 7–14 m range effectively released accumulated elastic energy. Post-relief stress peaks were significantly reduced and recovered more slowly. In terms of roadway stability, anchor-cable stress remained lower and more stable than under conventional drilling, with roadway side convergence and roof–floor convergence reduced by 63% and 51%, respectively. A comprehensive mechanical drilling–based anti-burst technology and equipment system was developed and successfully applied in engineering practice. These findings provide theoretical support and practical guidance for pressure relief and support strategies in hard coal seams of the Ordos region and similar mining conditions.

1 Introduction

Rock bursts are common dynamic disasters in underground coal mining operations (PAN et al., 2023; QI et al., 2019; Linming et al., 2022), and their occurrence in the Ordos region has been steadily rising in recent years. Some mines in the area experience severe roadway deformation and dynamic phenomena, which hinder safe production and limit high-quality capacity release. Consequently, research on rock bursts in the Ordos region has become a key focus in both academic and engineering fields (Yixin et al., 2020; HAN et al., 2022; GUO et al., 2021; Shankun, 2021).

Hard coal seams are a major geological factor contributing to rock bursts in the region because they enable significant elastic energy to accumulate around roadways, resulting in pronounced dynamic effects. About 86.8% of rock bursts occur along roadways (Jun et al., 2024; Jun et al., 2022; Zhang S. et al., 2025). Researchers have extensively studied the challenges of controlling the surrounding rocks around roadways within hard coal seams. Shankun (2021) identified hard roofs and hard coal seams as critical geological features of rock bursts, with coal body sliding in the roadway ribs as the primary failure mode. Pan et al. (2020) developed a three-tier support theory and technical system to prevent rock bursts in coal mine roadways. Additionally, Pan et al. (2024) analyzed and summarized the “three highs and one low” engineering challenges in deep rock burst prevention and proposed targeted localized pressure relief techniques. Liu et al. (2024) introduced the “stress differential gradient-induced rock burst mechanism” and recommended a collaborative approach involving active and passive support along with coal seam modification. Zhu et al. (2015) optimized support techniques and pressure relief parameters for severe deformation in gob-side roadways. Wu et al. (2021) proposed a “pressure relief-support-protection” strategy that effectively reduced deformation in deep roadway surrounding rocks. Gao et al. (2020) developed a “strong-weak-strong” structural model and combined “internal support external pressure relief” techniques for rock burst prevention. Shengrong et al. (2022) introduced a synergistic “external anchoring-internal pressure relief” control technology for large cross-section roadways to mitigate significant deformation. Kang (2024) advanced hydraulic fracturing, high-toughness bolts, and other surrounding rock control technologies, enhancing roadway stability. Yunliang et al. (2020) established the “pressure relief-stabilization” theory to improve energy absorption in anchoring systems. Weng and Yinghao (2019) proposed an integrated “drilling-cutting-pressuring” technology that lowered rock burst risks in gob-side roadways.

Globally, significant progress has been made in understanding the mechanisms, prediction, and control of rock bursts, especially in deep and high-stress mining environments. Dai et al., (2025) Studies in various countries have focused on developing constitutive models to simulate rock burst behavior, analyzing fault-slip seismicity influences, and proposing empirical methods for seismic hazard assessment. These efforts have deepened the theoretical understanding of stress redistribution, the identification of burst-prone zones, and energy release mechanisms (Zhang et al., 2026). Such insights provide valuable perspectives for improving rock burst prevention strategies under complex geological conditions, including those in China’s hard coal seams.

In summary, the academic and engineering communities mainly adopt two approaches to manage surrounding rock in roadways: “pressure relief” and “support.” The “pressure relief” approach centers on reducing the strength of surrounding rock to facilitate the release of stored elastic energy, while the “support” method focuses on reinforcing the surrounding rock near the roadway to improve resistance to rock bursts (Zhilun et al., 2012; Zhang L. et al., 2025).

For mines prone to rock bursts, maintaining overall roadway support strength while applying high-intensity pressure relief presents a critical challenge and a key area of focus. To address the pressure relief-support conflict in hard coal seams and develop effective control methods in the Ordos region, this study systematically examined the coupled mechanism of pressure relief and support for rock burst prevention through theoretical analysis and numerical simulation (Wu et al., 2025). Based on these findings, a mechanical reaming-based pressure relief technology for hard coal seams was proposed and implemented at the Menkeqing Coal Mine. The effects of pressure relief and support were evaluated using multiple methods to validate the technology’s effectiveness in preventing rock bursts. These findings offer both theoretical and practical guidance for similar conditions in the Ordos mining region.

2 Characteristics of pressure relief by drilling in hard coal seams

The uniaxial compressive strength (UCS) is a crucial measure of coal hardness. Reports on the impact potential of the main mining coal seams were gathered from 40 coal mines across key rock-burst-prone regions in China, including East Shandong, Central Shandong, West Shandong, Yima in Henan, Binchang in Shaanxi, and Ordos. Statistical analysis of the UCS values summarized in Table 1 and Figure 1 shows that the UCS of coal seams in the Ordos area generally ranges from 25 to 35 MPa, which is higher than in Shandong, Henan, and Shaanxi. This suggests that the coal seams in the Ordos region are the hardest among the major rock burst-prone mining districts in China.

Table 1

Table 1. Statistical data on uniaxial compressive strength of coal seams in major rock burst-prone mining areas in China.

Figure 1

Figure 1. Statistical analysis of uniaxial compressive strength of coal seams in major rock burst mining areas in China.

Drilling for pressure relief is the most common method used in the Ordos region, following the national standard Impact Ground Pressure Measurement, Monitoring, and Prevention Methods (GB/T25217.10). Boreholes here typically have diameters of 150–200 mm and depths of 15–25 m. Under hard coal seam conditions, this approach has distinct features.

1. Boreholes outside the advance influence zone are structurally stable. Beyond 70 m ahead of the working face, they rarely collapse and generally remain intact (Figure 2a).

2. Shear failure occurs when entering the advanced influence zone. As shown in Figure 2b, transverse splitting appears about 1 m inside the borehole opening, leading to delamination between the borehole and coal near the roadway. This shear-dominated failure mode suggests that the coal, under concentrated horizontal stress, tends to develop tensile-shear cracks along the borehole wall, weakening the shallow coal and reducing its ability to transfer load to the support system.

3. Borehole collapse causes significant damage to roadway surfaces. Figure 2c shows that once inside the influence zone, borehole deformation and sudden collapse often occur, with intense bulging of the sidewalls. These failure patterns indicate that stress around the borehole exceeds the coal’s peak strength, causing rapid plastic deformation and stress redistribution, which weaken nearby support structures. The ongoing collapse also points to a lack of effective pressure dissipation, highlighting limitations of standard drilling in high-strength coal seams.

4. Repeated pressure relief is needed due to the high strength of coal, which makes effective pressure management difficult. Mines often use high-density multi-round drilling with large-diameter boreholes. However, this repeated process can weaken roadway support structures (Figure 2d).

Figure 2

Figure 2. Characteristics of drilling pressure relief in hard coal seams. (a) Well-formed boreholes (b) Shear failure. (c) Borehole collapse (d) Repeated pressure relief.

In summary, although pressure relief drilling is widely adopted in the Ordos region, its effectiveness in hard coal seams is often suboptimal. Enhancing the efficiency of pressure relief drilling while minimizing damage to basic support structures remains a critical challenge in rock burst prevention research.

3 Coupled pressure relief and support mechanism for rock burst prevention in hard coal seams

3.1 Analytical solution for the radius of the plastic zone around boreholes

A mechanical model of a borehole under anisotropic in situ stress was established using classical elasticity (Figure 3). The model assumes unequal biaxial far-field loading, simulating the stress asymmetry typical of deep hard-coal seams. As shown in Figure 3, vertical and horizontal principal stresses generate an elliptical redistribution of stresses around the borehole. The resulting tangential stress concentrations along the borehole wall indicate where plastic deformation is likely to initiate. This model underpins the subsequent derivation of the plastic-zone radius and motivates enlarging boreholes in high-stress zones to shift the plastic boundary outward and more effectively release stored elastic energy.

Figure 3

Figure 3. Mechanical model of borehole under non-isobaric stress field conditions.

The stress around the borehole in polar coordinates is expressed by Equation 1, where the stress at any point within the model is determined by r and θ.

$\left\{\begin{array}{l}{\sigma }_{\mathrm{r}}=\frac{1}{2}P\left[\left(1+k\right)\left(1-\frac{R_0^2}{r^2}\right)+\left(1-k\right)\times \left(1-4\frac{R_0^2}{r^2}+3\frac{R_0^4}{r^4}\right)\cos 2\theta \right]\\ {\sigma }_{\theta }=\frac{1}{2}P\left[\left(1+k\right)\left(1+\frac{R_0^2}{r^2}\right)-\left(1-k\right)\times \left(1+3\frac{R_0^4}{r^4}\right)\cos 2\theta \right]\\ {\tau }_{\mathrm{r}\theta }=\frac{1}{2}P\left[-\left(1-k\right)\times \left(1+2\frac{R_0^2}{r^2}-3\frac{R_0^4}{r^4}\right)\sin 2\theta \right]\end{array}\right.(1)$

Where p represents the initial ground stress, K represents the lateral pressure coefficient for the coal body location, R₀ represents the borehole radius, σ_r represents the radial stress at point (r, θ) in the coal body, σ_θ represents the tangential stress at point (r, θ) in the coal body, and τ_rθ represents the shear stress at point (r, θ) in the coal body.

Substituting θ = 0° into Equation 1 can yield the tangential stress along the horizontal axis of the elastic zone of the surrounding rock.

${\sigma }_{\theta }=\frac{P_0}{2}\left[2+\left(k+1\right)\frac{{R_0}^2}{r^2}+3\left(1-\mathrm{k}\right)\frac{{R_0}^4}{r^4}\right](2)$

When k≠1, the total load on the horizontal axis of the elastic surrounding rock is given by

$Q^e=\underset{L\to \infty }{\lim }{\int }_{R_0}^L{\sigma }_{\theta }dr=PL(3)$

When plastic deformation occurs around a borehole, the stress distribution exhibits both elastic and plastic characteristics. For k ≤ 1, the tangential stresses in the elastic-plastic and elastic states along θ = 0° are shown in Figure 4.

Figure 4

Figure 4. Stress analysis diagram of the surrounding rock (k ≤ 1).

The total load on the horizontal axis of the surrounding rock can be expressed as

$Q^P={\int }_{R_0}^{R_s}{\sigma }_{\theta \left(r\le R_s\right)}dr+\underset{L\to \infty }{\lim }{\int }_{R_s}^L{\sigma }_{\theta \left(r>R_s\right)}dr(4)$

where ${{{{\sigma }_{\theta }}_{(}}_{r\le R_s}}_{)}$ and ${{{{\sigma }_{\theta }}_{(}}_{r>R_s}}_{)}$ represent tangential stresses in the plastic and elastic zones around the borehole on the horizontal axis, respectively (MPa).

Studies (Savin, 1960) have indicated that under non-isobaric conditions, the stress field along the horizontal axis within the plastic zone remains uniform.

${\sigma }_{\theta \left(r\le R_s\right)}=M\left(1-A\right){\left(\frac{r}{R_0}\right)}^{A-1}(5)$

The tangential stress ${{{{\sigma }_{\theta }}_{(}}_{r\le R_s}}_{)}$ in the elastic zone of the surrounding rock, denoted as σ_θ, was determined using Equation 2, as follows:

${\sigma }_{\theta \left(r>R_s\right)}=\frac{P}{2}\left[2+\left(k+1\right)\frac{R_{s1}^2}{r^2}x+3\left(1-\mathrm{k}\right)\frac{R_{s1}^4}{r^4}\right](6)$

The stress at the elastic-plastic interface should satisfy the yield condition, with the shear stress of the surrounding rock along the horizontal axis being zero. Substituting r = R_s1 into Equation 6 and solving it in conjunction with the M-C criterion equation can yield the following result:

${\sigma }_r{|}_{r=R_s}=\frac{\left(3k-1\right)P-B}{A+1}(7)$

To describe the σ_θ-r relationship in the surrounding rock in the elastic zone, Equation 6 is rewritten as

${\sigma }_{\theta \left(r>R_s\right)}=P+{\sigma }_{\theta }·F\left(r\right)(8)$

where F(r) is a function of r. The physical meaning of σ_θ is shown in Figure 4, which represents the stress difference between the elastic and plastic interface stresses and P of the surrounding rock on the horizontal axis (MPa).

By combining Figure 4 with the given equations, when r = R_s1, F(r) = 1. By substituting r = R_s1 into Equation 6 and F(r) = 1 into Equation 8, the two equations can be solved simultaneously to produce the following Equation 9:

${\sigma }_{\theta }=\left(2-k\right)P-{\sigma }_r{|}_{r=R_s}=\left(2-k-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)P(9)$

By combining Equations 6–8, the expression of F(r) can be obtained as Equation 10

$\begin{array}{l}F\left(r\right)=\left(\frac{1+k}{2\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)}-\frac{{\sigma }_r{|}_{r=R_s}}{\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)P}\right)\frac{R_{s1}^2}{r^2}\\ +\frac{3\left(1-k\right)}{2\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P_0}\right)}\frac{R_{s1}^4}{r^4}\end{array}(10)$

Hence, the tangential stress in the elastic zone along θ = 0° can be expressed as

${\sigma }_{\theta \left(r>R_s\right)}=P+\left({\sigma }_r{|}_{r=R_s}-P\right)\left[\begin{array}{l}\left(\frac{1+k}{2\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)}-\frac{{\sigma }_r{|}_{r=R_s}}{\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)P}\right)\frac{R_0^2}{r^2}\\ -\frac{3\left(1-k\right)}{\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)}\frac{R_{s1}^4}{r^4}\end{array}\right](11)$

where σ_θ|r=Rs is the tangential stress on the vertical axis of the surrounding rock at the elastic-plastic interface (MPa). Its expression can also be solved under uniform stress field conditions as follows:

${\sigma }_r{|}_{r=R_s}=M\left(1-A\right){\left(\frac{R_s}{R_0}\right)}^{A-1}(12)$

By substituting Equations 5, 11 and 12 into Equation 4, the total load on the elastic-plastic surrounding rock along the vertical axis was determined. Applying the principle of total load conservation and solving for the elastic surrounding rock total load, Equation 13 provides an approximate solution for R_s1 when k ≤ 1:

$R_{s1}=R_0{\left[\frac{\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)\left(P-M\right)+\left(k-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)\left(P-M\right)}{-M\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)-AM\left(1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)}\right]}^{\frac{1}{A-1}}(13)$

When k > 1 and plastic deformation occurred around the borehole (Figure 5), the radius of the plastic zone along the vertical axis R_s2 was nonzero. The tangential stress of the surrounding rock along the vertical axis exhibited both elastic and plastic stress states. By applying the conservation of the total load along the vertical axis and using the plastic zone stress formula under uniform stress field conditions, R_s2 can be determined as Equation 14.

${R_s}_2=R_0{\left[\frac{\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)\left(kP-M\right)+\left(k-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)\left(kP-M\right)}{-M\left(2k-1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)-AM\left(1-\frac{{\sigma }_r{|}_{r=R_s}}{P}\right)}\right]}^{\frac{1}{A-1}}(14)$

Figure 5

Figure 5. Stress analysis diagram of the surrounding rock on the vertical axis (k > 1).

3.2 Analysis of factors influencing borehole pressure relief effectiveness

The surrounding rock was assumed to transition continuously into an ideal plastic state, with the Mohr envelope for plastic deformation represented as a straight line defined by cohesion c and internal friction angle φ (Figure 6).

Figure 6

Figure 6. Maximum shear stress strength criterion.

From the Mohr stress circle in the ultimate state, the plastic condition can be derived as follows:

$\sin \varphi =\frac{{\sigma }_1^{\prime }-{\sigma }_3^{\prime }}{{\sigma }_1^{\prime }+{\sigma }_3^{\prime }+2x}=\frac{\sqrt{{\left({\sigma }_r-{\sigma }_{\theta }\right)}^2+4{\tau }_{r\theta }^2}}{{\sigma }_r+{\sigma }_{\theta }+2c\cot \varphi }(15)$

Substituting Equation 1 into the plastic condition in Equation 15 can yield an implicit equation for the plastic zone boundary of the borehole surrounding rock as a function of r and θ:

$\begin{array}{l}f\left(r,\theta \right)={\left[P\left(k+1\right)\frac{R_0^2}{r^2}-P\left(k-1\right)\cos 2\theta \left(1+3\frac{R_0^4}{r^4}-2\frac{R_0^2}{r^2}\right)\right]}^2\\ +{\left[P\left(k-1\right)\sin 2\theta \left(1+3\frac{R_0^4}{r^4}+2\frac{R_0^2}{r^2}\right)\right]}^2\\ -\frac{\left(1-\cos 2\varphi \right)}{2}\left\{{\left[P\left(k+1\right)\frac{R_0^2}{r^2}-2\cos 2\theta ·P\left(k-1\right)\frac{R_0^2}{r^2}\right]}^2-4c\right\}\\ -2c\sin 2\varphi \left[P\left(k+1\right)-2\cos 2\theta ·P\left(k-1\right)\frac{R_0^2}{r^2}\right]-4c^2\end{array}(16)$

When f(r,θ) = 0, the boundary line separating the elastic and plastic zones of the rock surrounding the borehole could be determined. As shown in Equation 16, the factors influencing the plastic zone include the lateral pressure coefficient, coal body strength, and borehole diameter. For a specific roadway, the lateral pressure coefficient is assumed to remain constant, allowing the analysis to focus on the effects of coal body strength and borehole diameter on the extent of the plastic zone in Table 2.

Table 2

Table 2. Summary of mechanical parameters for different coal bodies.

3.2.1 Influence of coal body strength

Mechanical parameter testing—including uniaxial compression and Brazilian (indirect tensile) splitting—was performed on coal specimens spanning a range of strengths to determine cohesion (c), internal friction angle (φ), uniaxial compressive strength (UCS), and shear-strength parameters for six coal types. Under identical in situ geological conditions for all six coal bodies, Equation 16 was applied to compute the plastic-zone radius of each borehole, thereby quantifying the relationship between coal strength characteristics and the extent of the plastic zone (Jia et al., 2024).

As shown in Figures 7, 8, the radius of the plastic zone around the borehole is negatively correlated with the coal’s uniaxial compressive strength (UCS) and shear strength, indicating that higher strength corresponds to a smaller plastic-zone radius and greater resistance of the surrounding rock to damage. Xu et al., (2022) when the UCS exceeds 13.69 MPa, the slope of the curve decreases markedly, implying diminishing sensitivity of the plastic-zone radius to further increases in strength. For hard coal, a smaller plastic-zone radius enhances roadway stability; however, the greater capacity for elastic-energy accumulation means that failure may release more energy and manifest as more severe rock bursts.

Figure 7

Figure 7. Relationship between uniaxial compressive strength and plastic zone.

Figure 8

Figure 8. Relationship between coal strength and plastic zone radius.

3.2.2 Influence of borehole diameter

By setting P = 16 MPa, k = 0.4, c = 2 MPa, and φ = 25°, the relationship between the borehole diameter and plastic zone of the surrounding rock was analyzed by varying R₀. The results are shown in Figure 9.

Figure 9

Figure 9. Relationship between the diameter of the pressure relief hole and plastic zone.

Figure 9 illustrates that as the borehole diameter increased, the radius of the plastic zone also gradually increased, while the developmental morphology of the plastic zone remained unchanged. The radial stress curves for boreholes with the diameters R₀ = 0.2, 0.3, and 0.4 m were extracted and analyzed (Figure 10).

Figure 10

Figure 10. Relationship between the diameter of the pressure relief hole and the plastic zone.

Figure 10 shows that increasing the borehole diameter from 0.20 to 0.30 m enlarges the plastic-zone radius by approximately 0.43 m, whereas a further increase from 0.30 to 0.40 m yields only about 0.26 m. This behavior indicates diminishing returns in the pressure-relief effect with continued diameter enlargement; therefore, simply enlarging the borehole diameter cannot indefinitely improve pressure relief.

The analysis further shows that, as the coal strength increases, stress-induced cracking around the borehole is reduced, making the borehole less susceptible to damage and decreasing the effectiveness of pressure relief. Within practical limits, increasing the borehole diameter remains the most direct way to enhance pressure relief; however, excessive diameters may interfere with bolt-and-cable support systems in roadways and thereby compromise control of the surrounding rock.

3.3 Coupled pressure relief and support method for hard coal seams

Maintaining roadway stability requires dissipating the accumulated elastic energy in the coal while preserving the load-bearing capacity of the support system—i.e., achieving a balance between pressure relief and support. To this end, a coupled technology termed “shallow support, deep pressure relief” was developed specifically for hard coal seams (Figure 11).

1. Shallow support. Treating the roadway as a circular borehole, the surrounding coal is partitioned into three zones: a fractured zone, a plastic-softening zone, and an elastic zone. In the fractured and plastic-softening zones beyond the peak-stress location, support is prioritized without increasing borehole diameter or drilling density (i.e., without reducing spacing). Conventional large-diameter shallow boreholes are retained in these zones to ensure that the shallow support members of the roadway remain undamaged.

2. Deep pressure relief. Within the peak-stress region, the interior elastic core is pressure-relieved by enlarging the borehole diameter. This enlargement drives a larger portion of coal to transform from elastic to plastic behavior, thereby releasing substantial elastic energy. As a result, the total elastic energy stored in the elastic core of hard coal is reduced, achieving effective pressure relief and impact (rock-burst) mitigation.

Figure 11

Figure 11. Coupled unloading and support technology for expansion borehole pressure relief in hard coal seams.

Combined strategy. Coupling shallow small-diameter borehole support with deep borehole enlargement for pressure relief lowers the stress concentration at the original peak and shifts the peak deeper into intact coal—producing a “peak-shaving and pressure-transfer” effect. Meanwhile, preserving the existing support capacity of the surrounding rock minimizes roadway deformation and enhances overall impact resistance.

4 Numerical simulation of coupled pressure relief and support in hard coal seams

4.1 Mechanism of coupled pressure relief and support through borehole enlargement

To elucidate the mechanism of the coupled pressure-relief–and-support method in hard-coal seams and to assess the effects of key parameters—borehole diameter, enlargement diameter, and enlargement position—on impact-prevention performance, a case study was conducted under the engineering–geological conditions of the Menkeqing 3106 working face. A stratified (layered) numerical model was constructed from the measured physical and mechanical properties of each rock stratum. Using FLAC^3D, we simulated the evolution of the stress and displacement fields of the surrounding rock during pressure relief. The three-dimensional model reflecting the in situ operating conditions is shown in Figure 12.

Figure 12

Figure 12. Numerical model structure.

The model dimensions were length (Y) × width (X) × height (Z) = 45 m × 8 m × 15 m. Boreholes were drilled along the positive Y-axis. The Mohr-Coulomb plastic strength criterion was applied, and the coal seam and surrounding rock were simulated using brick elements.

The boundary conditions for the model were as follows: (1) the horizontal movement was restricted to the sides; (2) the vertical movement was restricted at the bottom; (3) a vertical stress of 17 MPa was applied at the top to reflect the burial depth of the coal seam; and (4) a lateral pressure coefficient of 0.8 determined through in situ stress testing. The physical and mechanical parameters of the coal and rock for the simulations were obtained from laboratory tests of the surrounding rock properties, and the detailed parameters are listed in Table 3.

Table 3

Table 3. Physical and mechanical parameters of coal rock.

After achieving the initial equilibrium in the model, simulations were conducted to analyze the stress and deformation patterns of the surrounding rock under three conditions: (1) roadway excavation, (2) conventional borehole drilling, and (3) coupled pressure relief and support through borehole enlargement. The roadway dimensions were 5 m × 5 m, and cable anchor units were installed after excavation. The anchors measured Φ22 × 2500 mm, with full-length anchoring and a row spacing of 1 m × 1 m. Each row included 12 anchor units with an initial anchoring force of 60 kN. Following anchor installation, simulations were performed for conventional borehole drilling and coupled pressure relief and support. The conventional drilling diameter was 150 mm, the enlargement diameter was 350 mm, the enlargement point was 6 m from the coal wall, and the drilling spacing was 1 m. The vertical stress distribution and plastic zone expansion cloud diagrams for the coal seam in the three scenarios are shown in Figure 13.

Figure 13

Figure 13. Comparison of vertical stress distribution and plastic zone cloud maps in coal seams under different drilling methods. (a) Comparison of vertical stress distribution cloud maps. (b) Comparison of plastic zone expansion cloud maps.

The lateral pressure coefficient was set to 0.8 based on in situ stress measurement data obtained from the Menkeqing Coal Mine, which is located in the Ordos basin, where the horizontal stress is typically lower than the vertical stress due to tectonic stability and burial depth characteristics. Field tests using hydraulic fracturing and overcoring methods provided a consistent range of 0.75–0.85, supporting the use of 0.8 as a representative average.

The initial anchoring force of 60 kN was selected based on industry practice and site-specific roadway support designs in hard coal seams, as recommended by national standards for deep coal seam anchorage in dynamic pressure zones. In-situ pull-out tests also confirmed that this pre-tension level ensures anchorage efficiency while preventing overstressing of bolts in high-deformation zones.

Figure 13a shows that roadway excavation produces an “elastic core” about 6 m inside the coal wall due to stress concentration. Conventional pressure-relief drilling reduces the size of this elastic core, but the sector between the borehole and the roof exhibits only a slight decrease in stress concentration. In contrast, the coupled pressure-relief with borehole enlargement nearly eliminates the elastic core and markedly lowers the stress concentration. As shown in Figure 13b, the coupled scheme substantially expands the plastic zone between adjacent boreholes, thereby enhancing pressure relief. However, within roughly 1 m of the coal wall the plastic zone changes little relative to conventional drilling, indicating that the coal wall retains some load-bearing capacity. This outcome demonstrates that the coupled pressure-relief method minimizes disturbance to the roadway’s shallow support system.

4.2 Influence of borehole diameter on the coupled support and pressure relief effect

Numerical simulations indicate that, within the coupled pressure-relief–and-support scheme with borehole enlargement, the shallow-borehole diameter primarily governs roadway support performance, whereas the deep-borehole diameter mainly controls pressure relief within the coal mass. To quantify these effects, two one-factor-at-a-time (controlled-variable) simulation sets were performed: (1) the deep-borehole diameter was fixed at 350 mm while the shallow-borehole diameter varied from 100 to 200 mm; and (2) the shallow-borehole diameter was fixed at 150 mm while the deep-borehole diameter varied from 200 to 400 mm. Figure 14 presents the deformation curves of the surrounding rock and the effective constraint (restraining) force mobilized by the anchor bolts for different shallow-borehole diameters, whereas Figure 15 shows the vertical-stress profiles and peak-stress variation for different deep-borehole diameters.

Figure 14

Figure 14. Influence of shallow drilling diameter on support effectiveness. (a) Surrounding rock deformation curves (b) Effective anchor bolt constraint force curves.

Figure 15

Figure 15. Influence of deep drilling diameter on pressure relief effect. (a) Vertical stress variation curves (b) Peak stress variation curves.

Figure 14a shows that, with increasing shallow-borehole diameter, roof–floor convergence increases monotonically, whereas sidewall convergence first decreases and then increases. For diameters of 100–130 mm, sidewall deformation is slightly lower than the baseline (no pressure relief) because drilling alleviates stress concentrations in the wall coal and reduces surrounding-rock deformation. When the shallow-borehole diameter exceeds 130 mm, sidewall deformation rises sharply, reaching 429.3 mm at 200 mm—a 49.3% increase relative to the baseline. Figure 14b further indicates that the effective constraining force mobilized in the anchor bolts along the sidewalls and roof declines as the shallow-borehole diameter increases, with a steeper rate of reduction at larger diameters. Collectively, these results suggest that diameters >130 mm begin to degrade the shallow support system, undermining effective control of roadway deformation.

Figure 15a illustrates that, as the deep-borehole diameter increases, the vertical-stress distribution within the coal body consistently exhibits a double-peak pattern. The shallow-side peak stress gradually decreases, whereas the deep-side peak stress increases progressively. Figure 15b shows that when the deep-borehole diameter increases from 200 to 400 mm, the shallow-side peak stress decreases from 14.48 to 9.31 MPa (a reduction of 5.17 MPa), while the deep-side peak stress increases from 8.42 to 9.72 MPa (an increment of 1.30 MPa). These results indicate that enlarging the deep-borehole diameter enhances the pressure-relief efficiency, redistributing the concentrated stress deeper into the coal mass.

It is noteworthy that the effectiveness of both shallow and deep borehole diameters is influenced by other model parameters, including the lateral pressure coefficient and bolt pretension. Additional comparative simulations with lateral pressure coefficients ranging from 0.6 to 1.0 revealed that higher coefficients produce a narrower plastic zone and smaller sidewall deformation due to greater lateral confinement, whereas lower values intensify stress concentration near the roadway. Likewise, simulations with bolt pretension levels between 40 and 80 kN showed that increasing pretension effectively suppresses roof–floor convergence, though the improvement becomes marginal beyond 70 kN.

These findings highlight the necessity of balanced parameter selection—optimizing borehole geometry and support parameters to prevent overdesign while maintaining roadway stability under dynamic loading conditions.

5 Engineering practice of coupled support and pressure relief in hard coal seams

5.1 Engineering trial overview

The 3106 working face of the Menkeqing Coal Mine (side-exposed panel) primarily extracts the 3-1 seam, which shows a strong propensity for dynamic impacts. The uniaxial compressive strength reaches 33.54 MPa. During excavation and retreat of the 3106 return-air roadway, numerous high-energy events exceeding 10⁵ J were recorded, accompanied by frequent coal-cannon (spall) events and pronounced sidewall deformation, indicating a high risk of dynamic impacts. Conventional large-diameter boreholes did not provide effective pressure relief. To address this, an industrial trial of the coupled support and pressure-relief technology was implemented at the 3106 working face to enhance pressure relief.

Figure 16a shows that the trial site was arranged along the 3106 return-air roadway between 579 and 649 m from the retreat channel and about 700 m ahead of the working face to avoid the influence of advanced abutment pressure. In this region, no borehole pressure relief was applied during excavation; the support system remained intact, and the coal wall showed good integrity. Three parameter sets—A, B, and C—were established for comparative testing:

Figure 16

Figure 16. Engineering testing location and on-site situation. (a) Engineering trial location (b) Mechanical borehole enlargement tool (c) On-site pressure relief via borehole enlargement.

Area A (20 m): Mechanical borehole enlargement for pressure relief with 1 m spacing and 20 m depth. Sectional enlargement began at 6 m from the collar and was repeated every 3 m, leaving 1 m un-enlarged between segments. Shallow-borehole diameter = 150 mm; deep-borehole diameter = 350 mm.

Area B (20 m): Same as Area A except borehole spacing = 2 m; all other parameters unchanged.

Area C (30 m): Conventional large-diameter pressure-relief drilling with 2 m spacing, 20 m depth, and 150 mm diameter.

On site, hydraulically actuated mechanical under-reamers were used. The tools expanded under high hydraulic pressure and retracted under low pressure, enabling rearward rotary cutting to enlarge the borehole. Field implementation of borehole enlargement and pressure relief is shown in Figure 16.

The mechanical under-reaming tool employed a hydraulically actuated expansion mechanism controlled by adjustable water pressure. During on-site operations, the working pressure ranged from 18 to 25 MPa, and the flow rate was maintained at ≈80–100 L/min, ensuring reliable blade deployment and stable rearward rotary cutting. Each reaming head was fitted with tungsten-carbide (cemented-carbide) cutters. Operational wear testing indicated an average wear rate of ≈2.3 mm per 100 m of drilling under typical coal-seam hardness (UCS ≈30 MPa). Field practice further suggests that a standard cutter set can complete ≈400–500 m of borehole enlargement before replacement. These parameters provide practical guidance for hydraulic-system configuration and cutter-life planning in future applications.

5.2 Inspection of pressure relief effect via borehole enlargement

To assess the effectiveness of the mechanical borehole enlargement method in releasing accumulated elastic energy within the coal mass, the coal dust generated per meter during drilling and the number of “coal cannon” events per meter were recorded for Areas A, B, and C. The results are shown in Figures 17, 18.

Figure 17

Figure 17. Statistics of average drilling powder.

Figure 18

Figure 18. Statistics on the average number of “coal cannons” produced per meter.

As shown in Figure 17, the borehole enlargement section (6–20 m) generated significantly more coal dust than conventional drilling, peaking at 225 kg/m, which was 3.1 times that of conventional drilling in Area C. Compared with Area B (2 m spacing), Area A (1 m spacing) exhibited an approximately 12% higher average yield of coal powder, suggesting a more effective release of elastic strain energy due to denser drilling.

Figure 18 indicates that a depth of 7–14 m within the coal wall corresponded to the stress concentration zone. In this zone, conventional drilling (Area C) only showed a slight increase in “coal cannon” events in the 12–14 m section, while the mechanical reaming boreholes in Areas A and B produced multiple coal cannon occurrences throughout the entire 7–14 m range. The average number of coal cannon events per meter was highest in Area A, indicating a more comprehensive release of internal energy, while Area B exhibited slightly lower intensity but a more stable occurrence pattern.

To further evaluate the stability of the reaming pressure relief effects, a statistical analysis of data dispersion was conducted. The standard deviation of drilling cuttings per meter was 21.4 kg in Area A, 17.2 kg in Area B, and 12.9 kg in Area C. The coefficient of variation (CV) for “coal cannon” events was 22% in Area A and 18% in Area B. These results suggest that although Area A achieved the highest mean effect, its pressure relief results exhibited slightly higher variability among individual boreholes, likely due to local coal heterogeneities or incomplete reaming in a few holes.

Nevertheless, more than 85% of boreholes in Area A produced drilling powder and coal cannon events within one standard deviation of the mean, confirming that the pressure relief effects were overall consistent and reliable. The results also demonstrate that the high-intensity reaming did not cause instability or over-destressing in the surrounding coal.

To evaluate the static load relief effect of mechanical borehole enlargement, 21 stress gauges were installed on the primary wall of the 3106 return air roadway across three sets, with measurements performed at seven depths (2, 4, 6, 8, 10, 12, and 14 m). Additionally, six gauges were installed on the auxiliary wall at 2 and 3 m depths (Figure 19). Stress variations before and after drilling are shown in Figure 20.

Figure 19

Figure 19. Monitoring position of coal seam stress.

Figure 20

Figure 20. Stress monitoring curves of coal seams in different regions and depths. (a) Stress monitoring curves at different depths in Area A (b) Stress monitoring curves at different depths in Area B (c) Stress monitoring curves at different depths in Area C.

The monitoring results further support the effectiveness of mechanical reaming. In Areas A and B, coal mass stress remained stable during and after drilling, with no significant surges except for a temporary spike in Area B on November 8 due to unrelated roof blasting. In contrast, stress gauges in Area C at 6 m depth showed a gradual upward trend, and those at 8–12 m recorded irregular surges up to 3.8 MPa, indicating residual stress concentration.

From the perspective of engineering optimization, while Area A demonstrated slightly superior pressure relief (higher coal powder yield and more active energy release), the incremental improvement over Area B was relatively small compared to the doubling of drilling workload required for 1 m spacing. Therefore, Area B (2 m spacing) presents a more efficient configuration, achieving a favorable balance between pressure relief performance and construction efficiency. This provides valuable guidance for the design and optimization of drilling layouts in similar high-strength coal seam conditions.

5.3 Examination of surrounding rock control effectiveness

To evaluate the impact of different pressure relief methods on roadway support damage and surrounding rock deformation, five new Φ21.8 × 4300 mm anchor cables were installed on the roof and on two sides in each of the three test areas (A, B, and C). The pre-tension force of the anchor cables was set to no less than 150 kN. Stress meters with self-storage functionality were attached to the ends of the anchor cables and were capable of storing data for over 2 weeks with a monitoring frequency of once every 10 min. Additionally, the “cross-point method” was employed to observe the surrounding rock deformation in the test areas.

Nine surface displacement monitoring stations were installed in the 3106 return air roadway to monitor the displacement between the two sides and between the roof and floor before and after pressure relief drilling operations. The layout of the monitoring points is shown in Figure 21.

Figure 21

Figure 21. Schematic of the monitoring plan.

Figures 22a,b display the monitoring curves of anchor cable stress for the roof and right side of the roadway in the three test areas, revealing the following observations. (1) During drilling, the anchor cables in the mechanical reaming zones exhibited uniform stress distribution without sudden fluctuations. (2) After drilling, the roof anchor cable stress followed the order Area C > Area B > Area A. The absolute value and tension increment in the conventional drilling zone were higher than those in the mechanical reaming zones, indicating that mechanical reaming had a minimal impact on the roof support system stability. (3) During drilling, the anchor cable stress increased stepwise in both conventional and mechanical reaming zones. However, the tension increment in the conventional drilling zone reached 16 kN, which was approximately twice that in the mechanical reaming zone. Because the supported area was within 0–4.3 m from the roadway side, this suggested that mechanical reaming reduced the stress on support structures, enhancing the stability of the surface support system. (4) After drilling, the anchor cable stress in both zones continued to increase stepwise. Within 1 month of drilling completion, the side anchor cable stress followed the same trend (Area C > Area B > Area A), demonstrating that mechanical reaming improved the stress distribution in roadway support structures.

Figure 22

Figure 22. Anchor-force monitoring curve. (a) Roof anchor cable stress monitoring on the right side (b) Right-side anchor cable stress monitoring.

Figures 23a,b present the monitoring curves for the average movement of the two sides and the roof and floor in the three regions, based on the average values of the three groups of measuring points in each region. During drilling, the movement of the roof and both sides increased. However, the deformation of the two sides in the conventional drilling zone was significantly greater, with a maximum rate of 51 mm/d. In contrast, the mechanical reaming pressure relief zones reduced the maximum side movement by 63% and the roof and floor movement by 51%. These results demonstrated that the mechanical reaming pressure relief provided significantly better control of roadway surrounding rock deformation compared to the conventional large-diameter drilling pressure relief.

Figure 23

Figure 23. Monitoring curve for deformation of tunnel surrounding rock. (a) Monitoring of the average movement of the two sides (b) Monitoring of the average movement of the roof and floor.

In terms of support performance, roof and side cable-bolt stresses in Zone B were only marginally higher than in Zone A, with post-drilling differences within ∼2–3 kN. Displacement monitoring likewise indicated that Zone A achieved ∼8–10% lower sidewall convergence. However, given the ∼50% increase in drilling quantity required for Zone A, the incremental improvement in deformation control is moderate. Therefore, while Zone A performs slightly better in both static-stress reduction and deformation control, Zone B offers a more favorable trade-off between pressure-relief effect and construction efficiency, providing practical guidance for field optimization in similar geological settings.

6 Conclusion

1. A non-isobaric (anisotropic) in situ stress–field model for coal-seam drilling was developed for Ordos conditions, and the plastic-zone radius around a borehole was derived mechanically. The radius is governed primarily by the lateral pressure coefficient $K_0$, coal-seam strength, and borehole diameter. Increasing coal strength reduces the plastic-zone radius, whereas enlarging the borehole increases it but with diminishing returns—hence, simply increasing diameter cannot deliver unlimited pressure-relief gains.

2. To address poor pressure relief and support damage associated with high-density, multi-round drilling in hard coals of the Ordos region, we developed a coupled rock-burst–mitigation technology based on mechanical reaming. Following the principle of “shallow protection for support, deep pressure relief, and unloading–support balance,” the scheme deploys large-diameter deep boreholes to release stored elastic energy while using smaller-diameter shallow boreholes to preserve the bearing capacity of the support system—achieving an optimized balance between pressure relief and support.

3. A 70-m coupled pressure-relief–and-support industrial trial was implemented at the 3106 working face (Menkeqing Coal Mine) and comprehensively evaluated:

i. The reamed (enlarged) sections produced 3.1× more coal dust than conventional drilling, with concentrated “coal-cannon” events in the 7–14 m interval—evidence of effective release of accumulated elastic energy.

ii. Relative to conventional drilling, the peak-stress zone exhibited lower stress levels and slower recovery, indicating superior pressure relief in areas dominated by static loading.

iii. During reaming-based pressure relief, roof and side cable bolts maintained stable stress without axial-force loss or failure; post-construction cable-bolt stresses were lower than under conventional drilling.

4. Within the mechanical-reaming zone, the maximum sidewall movement decreased by 63% and roof–floor movement decreased by 51%, significantly outperforming conventional large-diameter drilling in controlling roadway deformation. These outcomes provide practical engineering guidance for mines operating under similar conditions in the Ordos Basin.

Data availability statement

The raw data supporting the conclusions of this article will be made available by the authors, without undue reservation.

Author contributions

HG Han: Conceptualization, Formal Analysis, Funding acquisition, Resources, Writing – original draft, Writing – review and editing. DL-M: Methodology, Validation, Writing – original draft, Writing – review and editing. XJ-H: Software, Supervision, Visualization, Writing – original draft, Writing – review and editing. ZQ: Investigation, Software, Validation, Writing – original draft, Writing – review and editing. SY-W: Formal Analysis, Investigation, Project administration, Writing – original draft, Writing – review and editing. SM-M: Investigation, Project administration, Software, Validation, Writing – original draft, Writing – review and editing. ML: Conceptualization, Data curation, Resources, Writing – original draft, Writing – review and editing.

Funding

The authors declare that no financial support was received for the research and/or publication of this article.

Conflict of interest

Authors HG, XJ-H, SY-W, SM-M and ML were employed by China Coal (Ordos City) Energy Technology Co Ltd.

Authors HG, XJ-H, SY-W, SM-M and ML were employed by China Coal Energy Research Institute Co. Ltd.

Author ZQ was employed by China Coal Energy Company Limited.

The remaining author declares that the research was conducted in the absence of any commercial or financial relationships that could be construed as a potential conflict of interest.

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