Influence of seam spacing and lateral pressure on deviatoric loading of the withdrawal roadway surrounding rock under the goaf

Abstract

Addressing asymmetric deformation, failure in the withdrawal roadway, and support crushing caused by overburden structure at final mining under the goaf. The stress evolution pattern of the surrounding rock in the withdrawal roadway and the dynamic characteristics of the maximum deviatoric stress zones (MDSZ) during this period were analyzed. The effect mechanism of the lateral pressure coefficient (LPC) and the coal seam spacing (CSS) on the stress distribution of the surrounding rock was explained.The LPC and CSS have a nonlinear effect on the rock stress redistribution of the withdrawal roadway. As the LPC increases, the peak abutment pressure on the non-mining side initially rises then stabilizes. The change in principal stress ahead of the advance working face (AWF) shows an “N”-shaped trend, and the maximum deviatoric stress (MDS) on the non-mining side first decreases and then increases. As the CSS increases, the peak abutment pressure and deviatoric stress on the non-mining side initially decrease and then increase. When the CSS is 75 m is comparable to that in the absence of an overlying goaf. The evolution model of the MDSZ in the retreat channel surrounding rock comprises four stages: low values at both sidewalls → value growth at both sidewalls → maximum value and zone at the goaf-side rib (G-SR)→ maximum value and zone at the solid-side rib (S-SR). A collaborative control scheme of “anchor cable reinforcement support on the S-SR” and “pre-splitting roof cutting on the G-SR” was proposed. Field application shows the sidewall displacement is about 150 mm, and the roof displacement is about 200 mm, which ensures the safe withdrawal of equipment. The research results serve as a valuable reference for similar engineering practices.

Introduction

Due to the significant capital investment in mining equipment, coal mine production generally operates a single production system. To maintain continuous operation, the most critical aspect is the relocation of the working face, which involves the safe and efficient removal of production equipment. Currently, there are typically two methods for equipment removal. One is to pre-excavate a withdrawal channel. The other is to unhook the hydraulic supports and allow the shearer to self-excavate a withdrawal space as the mining face approaches the final pit limit. Regardless of the withdrawal way used, it is essential to consider potential accidents such as roof caving and support crushing caused by unsupported large-section roadway withdrawal, as well as spontaneous combustion of coal left in mined-out areas¹, which can lead to equipment failure.

The retreat roadway does not undergo large deformations, which is indispensable for equipment withdrawal.Scholars have had beneficial contributions in aspects such as the movement characteristics and fracture positions² of the overlying strata of the withdrawal roadway in the case of single-coal - seam mining, the instability and control of the surrounding rock³ affected by the strong mining effect^4–6 and large-scale geological structures⁷, and the reasonable spacing^8–10 of double withdrawal roadways for accelerating equipment withdrawal. Analysis of the withdrawal roadway’s roof shows multiple fracture points. The best overburden structure is behind the hydraulic support. It is found that the main cause of the roadway instability is that abutment pressure increases exponentially as the width of the coal pillar decreases. Moreover, active and passive support methods¹¹, as well as cavity-making¹² and roof cutting for pressure relief^13–15, are adopted to effectively control the roadway stability.

Compared to a single coal seam retreat roadway, the one beneath the goaf is subjected to the influence of remaining coal pillars and the structure of the overlying goaf, resulting in damage to the immediate roof of the underlying coal seam and compromising its integrity.Scholars^16–18 studied the stability of the coal pillar for roadway protection at final mining in close-distance coal seam mining, the stress propagation pattern and failure depth of the floor under the overlying residual coal pillars, as well as overlying strata failure and its effect on stress redistribution and clarified the source mechanism of dynamic pressure affecting the retreat roadway under the goaf. Meanwhile, in the mining of nearby coal seams, a complex superimposed stress field is created, which will inevitably make controlling the surrounding rock much more difficult in the withdrawal roadway at final mining of the lower coal seams. Scholars’ research^19,20 shows that the stability of overlying residual coal pillars, the fracture position of the main roof, and the mining are the key factors and causes of instability leading to extensive dynamic stress at final mining. At the same time, scholar²¹ has optimized the reasonable location of the withdrawal roadway for close-distance coal seam mining.

The above achievements are highly significant for guiding the control research of the withdrawal roadway at final mining beneath the goaf. However, When the CSS varies, the level of influence differs accordingly. Additionally, different LPC impact the distribution of the initial stress field and the direction of the principal stress²². However, current research does not address the asymmetric failure mechanism of the withdrawal roadway, the eccentric load mode of surrounding rock stress, or the principles of asymmetric support at final mining under the combined effects of varying LPC and different CSS. Therefore, this paper focuses on studying and analyzing the evolution of abutment pressure, principal stress, and deviatoric stress in the surrounding rock of the withdrawal roadway under different LPC and CSS. It identifies the asymmetric failure mechanism of the roadway, the stress superposition and eccentric load modes, and key deformation control areas, providing scientific guidance for their management. Additionally, the study provides valuable insights for stabilizing the retreat roadway during the final mining phase under goaf-induced stresses.

Engineering background and feature analysis

This mine mainly mines 2# coal and 5# coal. Those coal seams have a angle varying between 3° and 6°, making them nearly horizontal. As shown in Fig. 1, the average thicknesses of the two coal seams are approximately 2.4 m and 2.3 m respectively. 5# coal is located below 2# coal, and the average burial depth is 355 m. With the mine production, 2# coal has been mined out. In the final mining stage of 5# coal, both the withdrawal roadway and the return airway are located underneath the goaf. The withdrawal roadway is 20 m away from the return airway, and a 20 m horizontal offset separates the edge of the overlying goaf from the return airway.

Fig. 1.

Roadway failure induced by geological drilling and final mining operations: a schematic overview.

During final mining periods of the mining face of 5# coal, the withdrawal roadway is influenced not only by the stress redisturbance of the goaf, but continuously and vigorously disturbed by the advanced abutment pressure at final mining. Accordingly, exposure to superimposed dynamic and static loading at final mining disturbs the withdrawal roadway, inducing asymmetric stress and large deformation in the surrounding rock on both sides. Therefore, the scientific control of the withdrawal roadway is contingent upon a thorough analysis of its stress conditions at the final mining.

At final mining, as protective coal pillar for the retreat roadway continuously shortens, the surrounding rock of the withdrawal channel is continuously affected by the advanced abutment pressure. This causes a loss of integrity and strength in the surrounding rock mass due to the formation of numerous cracks. The rock mass continuously swells and extrudes into the withdrawal channel, leading to varying degrees of large deformation in the mining side, non-mining side, and roof of the withdrawal channel, as shown in Fig. 1.

Theoretical analysis of the rotation and load increment of the principal stress

The deviatoric stress differs from the abutment pressure. It not only combines the three principal stresses but also, according to elastoplastic theory, influences the plastic deformation of the surrounding rock after shear. By deriving the stress distribution around the hole using elastic mechanics, the following stress expressions in polar coordinates are obtained obtained²³.

1

σ_r is radial-directed stress; σ_θ is the tangential stress; τ_rθ is the shear stress; h is the depth of the roadway; λ is the LPC; a is the tunnel radius; the polar coordinates of any point are set as (θ, R_θ).Simplified deviatoric stress formula for plane problems in elastic mechanics²³:

2

Therefore, introducing the deviatoric stress as an analytical index provides a more comprehensive representation of the stress state at a point compared to abutment pressure. Furthermore, the intensity of mining-induced influence significantly affects the evolution process of the MDSZ in the surrounding rock of the retreat roadway, as shown in Fig. 2.

Fig. 2.

Schematic diagram of stress component evolution and evolution pattern of the MDSZ affected by mining activities in mining engineering.

Simulation and analysis of the stress evolution law of the surrounding rock in the withdrawal channel at the final mining under the goaf

To investigate the superposition and evolution rules of stress in the withdrawal roadway under different LPC and various CSS, taking the withdrawal roadway of the 5106 working face as the engineering background, FLAC3D was employed to investigate the evolution of abutment pressure, principal stress deflection, and deviatoric stress in the surrounding rock of the withdrawal roadway under the 2102 goaf.

Simulation method and model establishment

FLAC3D was employed for modeling and analysis of the mine’s geological conditions. The model measures 180 m in the X-direction and 80 m in the Z-direction. The X-direction aligns with the strike of the working face, the Y-direction indicates the dip of the working face, and the Z-direction represents the vertical axis, as shown in Fig. 3. This model was configured using the Mohr-Coulomb constitutive relation, with displacement fixed on all boundaries except the top. An initial stress field is applied to the top surface, with a calculated vertical stress load of approximately 7.45 MPa. This setup simulates the actual engineering excavation steps: (1) excavation of the 2102 goaf, (2) excavation of the withdrawal channel and intake airway, (3) excavation of the 5106 working face to various distances from the withdrawal channel.

Fig. 3.

Numerical simulation diagram.

The increasing abutment pressure on the surrounding rock of the withdrawal roadway under the goaf

Based on Fig. 4, with a 20 m layer spacing, the abutment pressures on withdrawal roadway sides demonstrate localized loading enhancement as the working face advances.

Fig. 4.

Under different LPC with a layer spacing of Y = 20 m, the evolution law of the abutment pressure as the protective coal pillar for the withdrawal roadway is X m.

1. The abutment pressures on both sides of the withdrawal roadway exhibit local load-increasing characteristics. As the protective coal pillar for the retreat roadway is mined to 30 m, the peak abutment pressures (PAP) on the G-SR of the withdrawal roadway with LPC of 0.8, 1.2, and 1.8 are 26.4%, 16.3%, and 10.1% higher than those on the S-SR respectively. As protective coal pillar is mined to 20 m, the PAP on the G-SR of the withdrawal roadway with LPC of 0.8, 1.2, and 1.8 are 53.2%, 37.7%, and 34.3% higher than those on the S-SR respectively; As protective coal pillar is mined to 10 m, the PAP on the G-SR of the withdrawal roadway with LPC of 0.8, 1.2, and 1.8 are 30%, 30.5%, and 29.6% higher than those on the S-SR respectively. Which shows that during the process when the protective coal pillar is mined from 30 m to 20 m, the local load-increasing phenomenon of the abutment pressures on both sides of the withdrawal roadway under different LPC becomes more and more obvious; during the process when the protective coal pillar is mined from 20 m to 10 m, the PAP gradient at the withdrawal roadway ribs under different LPC begins to decline.

2. As the LPC increases, the PAP on the S-SR of the withdrawal roadway rises and stabilizes within a certain range. As the working face is mined, the PAP on the S-SR of the withdrawal roadway with a LPC of 0.8 are lower than those of the withdrawal roadway with LPC of 1.2 and 1.8. The difference between the former and the latter two also becomes larger as the mining progresses, with the difference ranging from 0.7→2.6 MPa (λ = 1.2) and 0.9→2.8 MPa (λ = 1.8). The difference between the abutment pressures on the S-SR of the withdrawal roadway with a LPC of 1.2 and that with a LPC of 1.8 during the mining process of the working face does not exceed 0.5 MPa.

Evolution law of the principal stress in surrounding rock of the withdrawal roadway under various LPC

To prevent the cross-influence of the withdrawal channel and the intake airway on the magnitude and direction of the principal stress ahead of the AWF, these features are not temporarily excavated in the numerical simulation. A measuring line runs along the center line of the coal seam ahead of the AWF to collect data on the principal stresses. Five typical measuring points, labeled A through E, are selected for analysis, with points B and D representing the centers of the original withdrawal channel and the original intake airway, respectively. For convenience in subsequent study, the angle between the maximum principal stress (MPS) direction line and the horizontal line is defined as the MPS direction angle (α). This angle is determined by the counter-clockwise rotation of the horizontal line to the MPS direction line (0° ≤ α ≤ 180°). as shown in Fig. 5.

Fig. 5.

The range of variation of the direction angle (α) of the MPS. (The same below, and no further elaboration is needed)

As shown in Figs. 6 and 7, the following laws can be obtained from the MPS deflection angles and stress magnitudes at different measuring points ahead of the AWF under different LPC at the final mining: (1) When the working face is mined to 32 m, 22 m, and 12 m from point B (the center point of the original withdrawal channel), with the increase of the LPC, the deflection angles of the MPS at points B, C (the center point of the original intake airway), and D all first decrease and then increase. As the working face is mined to 2 m from point B, this rule is not obvious. With the working face being mined to 32 m from point B, under the conditions of LPC of 0.8, 1.2, and 1.8, the deflection angles of the MPS at point B ahead of the AWF are 79°, 38°, and 120° respectively; the deflection angles of the MPS at point C ahead of the AWF are 83°, 20°, and 133° respectively; the deflection angles of the MPS at point D ahead of the AWF are 84°, 3°, and 139° respectively. When the working face is mined to 22 m from point B, under the conditions of LPC of 0.8, 1.2, and 1.8, the deflection angles of the MPS at point B ahead of the AWF are 65°, 57°, and 75° respectively; the deflection angles of the MPS at point C ahead of the AWF are 72°, 15°, and 128° respectively; the deflection angles of the MPS at point D ahead of the AWF are 76°, 2°, and 165° respectively. When the working face is mined to 12 m from point B, under the conditions of LPC of 0.8, 1.2, and 1.8, the deflection angles of the MPS at point B ahead of the AWF are 69°, 68°, and 69° respectively; the deflection angles of the MPS at point C ahead of the AWF are 64°, 43°, and 65° respectively; the deflection angles of the MPS at point D ahead of the AWF are 164°, 6°, and 63° respectively.

Fig. 6.

With a layer spacing of Y = 20 m, under different LPC, the magnitudes and directions of the principal stresses when the working face is 32 m and 22 m away from point B.

Fig. 7.

With a layer spacing of Y = 20 m, under different LPC, the magnitudes and directions of the principal stresses when the working face is 12 m and 2 m away from point B.

• (2)In the mining process, the changing trends of the magnitudes of the maximum and minimum principal stresses ahead of the AWF under different LPC are the same, showing a trend of first increasing, then decreasing, and then increasing again, approximately in an “N” shape.
• (3)In the process of the coal pillar being mined to 32 m, 22 m, and 12 m from point B, the changing trends of the differences between the magnitudes of the principal stresses at measuring points A–E ahead of the AWF under different LPC are the same, first decreasing and then increasing. This reflects that the area closer to the working face is affected by the mining activity of the working face, and the shear stress on the surrounding rock is relatively large; the area farther away is affected by the overlying goaf, and the shear stress on the surrounding rock is relatively small; the area even farther away is affected by the coal pillars in the overlying goaf, and the shear stress on the surrounding rock is relatively large. As the working face is mined to 32 m from point B, under the condition of a LPC of 0.8, the differences between the magnitudes of the principal stresses at measuring points A–E are 10.56 MPa, 2.57 MPa, 1.48 MPa, 0.92 MPa, and 4.89 MPa respectively; under the condition of a LPC of 1.2, the differences are 7.1 MPa, 1.71 MPa, 2.15 MPa, 3.73 MPa, and 4.07 MPa respectively; under the condition of a LPC of 1.8, the differences are 8.3 MPa, 1.31 MPa, 1.39 MPa, 3.83 MPa, and 9.35 MPa respectively. With the coal pillar being mined to 22 m from point B, under the condition of a LPC of 0.8, the differences between the magnitudes of the principal stresses at measuring points B–E are 5.26 MPa, 2.17 MPa, 1.32 MPa, and 5.05 MPa respectively; under the condition of a LPC of 1.2, the differences are 2.2 MPa, 1.56 MPa, 4.05 MPa, and 4.15 MPa respectively; under the condition of a LPC of 1.8, the differences are 2.81 MPa, 0.8 MPa, 3.84 MPa, and 9.74 MPa respectively. As the coal pillar is mined to 12 m from point B, under the condition of a LPC of 0.8, the differences between the magnitudes of the principal stresses at measuring points B–E are 11.6 MPa, 5.68 MPa, 2.44 MPa, and 4.89 MPa respectively; under the condition of a LPC of 1.2, the differences are 8.5 MPa, 3.48 MPa, 4.11 MPa, and 4.3 MPa respectively; in a case of a LPC of 1.8, the differences are 8.7 MPa, 2.08 MPa, 3.56 MPa, and 9.86 MPa respectively.

Evolution pattern of the deviatoric stress in the surrounding rock of the withdrawal roadway under varying LPC

As shown in Fig. 8, as the longwall face progresses, the overlying goaf is 20 m away from the underlying coal seam, and the deviatoric stress of the rock in the withdrawal roadway under different LPC shows the following patterns.

Fig. 8.

With a layer spacing of Y = 20 m, under different LPC, evolution of deviatoric stress in the surrounding rock of the withdrawal roadway.

1. The MDSZ of the surrounding rock in the withdrawal channel shows asymmetric loading characteristics. With a 30-meter-wide coal pillar in the retreat roadway, and a LPC of 0.8, the MDSZ in the G-SR is in the transition area between the G-SR and the floor, and that in the S-SR is in the transition areas between the S-SR and the roof and floor; for the withdrawal channel with a LPC of 1.2, the MDSZ in the G-SR is in the transition area between the G-SR and the floor, and that in the S-SR is in the transition area between the S-SR and the floor; for the withdrawal channel with a LPC of 1.8, the MDSZ in the G-SR is in the transition areas between the G-SR and the roof and floor, and that in the S-SR is in the transition area between the S-SR and the floor. With a 20-meter-wide coal pillar in the retreat roadway, and a LPC of 0.8, the MDSZ in the G-SR is in the transition areas between the G-SR and the roof and floor, showing an inclined strip distribution, and that in the S-SR is in the transition areas between the S-SR and the roof and floor; for the withdrawal channels with LPC of 1.2 and 1.8, the MDSZ in the G-SR is in the transition area between the G-SR and the floor, and that in the S-SR is in the transition areas between the S-SR and the roof and floor. With a 10-meter-wide coal pillar in the retreat roadway, and the three LPC, the MDSZ in the G-SR is in the transition areas of the G-SR, roof and floor; for the S-SR, the MDSZ with LPC of 0.8 and 1.8 are in the transition areas between the S-SR and the roof and floor, and that with a LPC of 1.2 is in the transition area between the S-SR and the roof.

2. As the LPC increases, the MDS on the S-SR of the withdrawal channel initially decreases and then increases. With a 30-meter-wide coal pillar in the retreat roadway, the MDS in the S-SR of the withdrawal channels with LPC of 0.8, 1.2, and 1.8 are 6.5 MPa, 5.14 MPa, and 5.84 MPa, respectively. The MDS on the S-SR of the withdrawal channel with a LPC of 0.8 is 1.36 MPa higher than that with a LPC of 1.2. Additionally, the MDS with a LPC of 1.8 is 0.7 MPa higher than with 1.2. With a 20-meter-wide coal pillar in the retreat roadway, the MDS on the S-SR of the withdrawal channels with LPC of 0.8, 1.2, and 1.8 are 7.87 MPa, 6.30 MPa, and 7.00 MPa, respectively. The MDS on the S-SR of the withdrawal channel with a LPC of 0.8 is 1.57 MPa higher than with a coefficient of 1.2, and the value with a coefficient of 1.8 is 0.7 MPa higher than with a coefficient of 1.2. With a 10-meter-wide coal pillar in the retreat roadway, the MDS on the S-SR of the withdrawal channels, with LPC of 0.8, 1.2, and 1.8, are 10.9 MPa, 9.82 MPa, and 10.2 MPa, respectively. The MDS on the S-SR of the withdrawal channel with a LPC of 0.8 is 1.08 MPa higher than that with a coefficient of 1.2, and the stress with a coefficient of 1.8 is 0.38 MPa higher than with 1.2. With a 0 m wide coal pillar in the retreat roadway, the MDS on the S-SR of the withdrawal channels with LPC of 0.8, 1.2, and 1.8 are 12.1 MPa, 10.6 MPa, and 11.2 MPa respectively.

Patterns of abutment pressure increase in the surrounding rock of withdrawal channels under different CSS

As shown in Fig. 9, when the LPC is 1.2, the abutment pressures on roadway ribs at different distances from the overlying goaf follow the following rules.

Fig. 9.

Under different CSS with λ = 1.2, the evolution law of the abutment pressure as the protective coal pillar for the withdrawal roadway is X m.

1. The abutment pressures on both sides of the withdrawal roadway show an increasing load characteristic. With a 30-meter-wide coal pillar in the retreat roadway, the PAP on the G-SR of the withdrawal roadway without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 19.8%, 16.3%, and 12.6% higher than those on the S-SR respectively. With a 20-meter-wide coal pillar in the retreat roadway, the PAP on the G-SR of the withdrawal roadway without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 43.1%, 37.7%, and 27% higher than those on the S-SR respectively. With a 10-meter-wide coal pillar in the retreat roadway, the PAP on the G-SR of the withdrawal roadway without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 26.6%, 30.5%, and 73.1% higher than those on the S-SR respectively.

2. Once the longwall face has progressed a specific distance, as the distance from the overlying goaf increases, the PAP on the S-SR of the withdrawal roadway first decreases and then increases. With a 30-meter-wide coal pillar in the retreat roadway, the PAP on the S-SR of the withdrawal roadway without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 16.2 MPa, 14.7 MPa, and 19.1 MPa respectively. The PAP on the S-SR of the withdrawal roadway without an overlying goaf and at a distance of 75 m from the overlying goaf are 1.5 MPa and 4.4 MPa higher than that at a distance of 20 m from the overlying goaf respectively. With a 20-meter-wide coal pillar in the retreat roadway, the PAP on the S-SR of the withdrawal roadway without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 19.5 MPa, 17.5 MPa, and 21.6 MPa respectively. The PAP on the S-SR of the withdrawal roadway without an overlying goaf and at a distance of 75 m from the overlying goaf are 2.0 MPa and 4.1 MPa higher than that at a distance of 20 m from the overlying goaf respectively. With a 10-meter-wide coal pillar in the retreat roadway, the PAP on the S-SR of the withdrawal roadway without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 25.9 MPa, 24.6 MPa, and 24.9 MPa respectively. The PAP on the S-SR of the withdrawal roadway without an overlying goaf and at a distance of 75 m from the overlying goaf are 1.3 MPa and 0.3 MPa higher than that at a distance of 20 m from the overlying goaf respectively. With a 0-meter-wide coal pillar in the retreat roadway, the PAP on the S-SR of the withdrawal roadway without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 29.4 MPa, 26.9 MPa, and 34.9 MPa respectively. The PAP on the S-SR of the withdrawal roadway without an overlying goaf and at a distance of 75 m from the overlying goaf are 2.5 MPa and 8.0 MPa higher than that at a distance of 20 m from the overlying goaf respectively.

Evolution law of the principal stress of the surrounding rock of the withdrawal roadway under varying CSS

The study examines the evolution laws of the magnitude and direction of the principal stress at various positions ahead of the AWF during each mining process, considering different distances from the overlying goaf. To prevent the cross-influence of the withdrawal roadway and the intake airway on the principal stress ahead of the AWF, these two features are temporarily not excavated in the numerical simulation. A measuring line is set along the center line of the coal seam ahead of the AWF to record data on the maximum and minimum principal stresses. Five measuring points, labeled from point A to point E, are selected for analysis. Points B and D correspond to the center points of the original withdrawal roadway and the original intake airway, respectively. The various advancing stages studied are illustrated in Figs. 10 and 11.

Fig. 10.

With λ = 1.2, under different CSS, the magnitudes and directions of the principal stresses as the working face is 32 m and 22 m away from point B.

Fig. 11.

With λ = 1.2, under different CSS, the magnitudes and directions of the principal stresses as the working face is 12 m and 2 m away from point B.

1. At various distances of the working face from the retreat roadway, the deflection angles of the MPS at points A, B, C, and D ahead of the AWF decrease progressively at 20 m and 75 m from the overlying goaf. The pattern of the deflection angles of the MPS ahead of the AWF without an overlying goaf is not obviously regular. With a 30-meter-wide coal pillar in the retreat roadway, the deflection angles of the MPS at points A, B, C, and D ahead of the AWF without an overlying goaf, at a distance of 20 m from the goaf, and at 75 m from the goaf are 77°, 32°, 169°, and 161°; 70°, 38°, 20°, and 3°; 82°, 68°, 45°, and 12° respectively. With a 20-meter-wide coal pillar in the retreat roadway, the deflection angles of the MPS at points A, B, C, and D ahead of the AWF without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 156°, 179°, and 180°; 57°, 15°, and 12°; 74°, 61°, and 29° respectively. With a 10-meter-wide coal pillar in the retreat roadway, the deflection angles of the MPS at points A, B, C, and D ahead of the AWF without an overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 70°, 46°, and 178°; 68°, 43°, and 6°; 72°, 64°, and 49° respectively. The deflection angles of the maximum principal stress at points A, B, C, and D ahead of the AWF, without an overlying goaf, at 20 m from the goaf, and at 75 m from the goaf are 67° and 79°; 64° and 24°; 66° and 57°, respectively.

2. As the coal pillar is mined to 32 m from point B, the differences between the principal stresses at measuring points A - E under the conditions of no overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 10.4 MPa, 1.4 MPa, 1.9 MPa, 2.37 MPa, and 1.98 MPa; 7.1 MPa, 1.71 MPa, 2.15 MPa, 3.73 MPa, and 4.07 MPa; 7.3 MPa, 2.2 MPa, 1.13 MPa, 1.24 MPa, and 1.77 MPa respectively. When the coal pillar is mined to 22 m from point B, the differences between the principal stresses at measuring points A - E under the conditions of no overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 3.6 MPa, 1.41 MPa, 2.12 MPa, and 2.16 MPa; 2.2 MPa, 1.56 MPa, 4.05 MPa, and 4.15 MPa; 3.7 MPa, 1.9 MPa, 1.02 MPa, 1.62 MPa, and 1.77 MPa respectively. With the coal pillar being mined to 12 m from point B, the differences between the principal stresses at measuring points A - E under the conditions of no overlying goaf, at a distance of 20 m from the overlying goaf, and at a distance of 75 m from the overlying goaf are 9.0 MPa, 3.1 MPa, 1.53 MPa, and 2.22 MPa; 8.5 MPa, 3.48 MPa, 4.11 MPa, and 4.3 MPa; 7.4 MPa, 4.1 MPa, 2.1 MPa, and 1.6 MPa respectively. At a standoff distance of 2 m between the working face and point B, the differences between the at measuring points A-E under the conditions of no overlying goaf, at a distance of 20 m from the overlying goaf, and a distance of 75 m from the overlying goaf are 7.5 MPa, 2.3 MPa, and 2.05 MPa; 8.6 MPa, 4.52 MPa, and 3.95 MPa; 7.5 MPa, 4.2 MPa, and 1.77 MPa, respectively.

Development mechanism of deviatoric stress in the retreat roadway surrounding rock with varying CSS

As can be seen from Fig. 12, with the advancement of the working face, development mechanism of deviatoric stress in the retreat roadway surrounding rock under different CSS follow the following rules.

Fig. 12.

With λ = 1.2, under different CSS, evolution of deviatoric stress in the surrounding rock of the withdrawal roadway.

1. With a 30-meter-wide coal pillar in the retreat roadway, the MDSZ on the G-SR is significantly higher than on the S-SR: 40% higher without overlying goaf, 10% higher when 20 m from the overlying goaf, and 11.9% higher at 75 m from the overlying goaf. At this point, the peak area of the stopping wall is in the transition zone between the stopping wall and the floor, while the peak area of the non-stopping wall is in the transition zone between the non-stopping wall and the floor. With a 20 m wide coal pillar in the retreat roadway, the MDSZ on the G-SR increases further: 72.1% higher than that of the non-overlying goaf, 62.5% higher than that of the overlying goaf at 20 m, and 31.6% higher than that of the overlying goaf at 75 m. Regarding the location of the peak areas, the G-SR remains in the transition zone between the G-SR and the floor, rather than the S-SR shifting to the transition zone between the S-SR and the roof and floor. With a 10-meter-wide coal pillar in the retreat roadway, differences in the MDSZ are as follows: 3.6% higher than the non-overlying goaf, 15.2% higher than the overlying goaf at 20 m, and 55.3% higher than the overlying goaf at 75 m. At this stage, the peak area of the G-SR extends into the transition zone between the G-SR and the roof and floor, with the regional scope expanding. The peak area of the S-SR is located at the boundary between the S-SR and the roof of the withdrawal channel.

2. With a 30-meter-wide coal pillar in the retreat roadway, the MDS in the no-overlying goaf, 20 m, and 75 m from the goaf are 5.67 MPa, 5.14 MPa, and 6.88 MPa, respectively. Compared to the 20 m distance, the stress in the area without goaf is 0.53 MPa lower, and at 75 m it is 1.74 MPa higher. With a 20-meter-wide coal pillar in the retreat roadway, the corresponding MDS are 6.69 MPa, 6.30 MPa, and 7.99 MPa. Compared to the 20 m mark, the stress in the no-goaf area is 0.39 MPa lower, and at 75 m it is 1.69 MPa lower. When mining reaches 10 m, the peak stresses are 10.4 MPa, 9.82 MPa, and 9.86 MPa. Compared with 20 m, the stress in the no-goaf zone increases by 0.53 MPa, while at 75 m, it increases by only 0.04 MPa. When mining reaches 0 m, the peak stresses are 6.69 MPa, 6.30 MPa, and 7.99 MPa. Compared to 10 m, the stress in the no-goaf area increases by 0.90 MPa, and at 75 m by 3.6 MPa.

In summary, during the transition from 30 m to 20 m, regardless of the distance from the overlying goaf, there is a gradual increase in the load peak in the deviatoric stress of the surrounding rock. In the process from 20 m to 10 m, the load-bearing capacity of the area without goaf and at 20 m from the goaf begins to weaken, whereas the loading at 75 m from the goaf becomes more pronounced.

Analysis of how various factors affect the stress evolution law of surrounding rock around the withdrawal channel under the goaf

By changing the two parameters of LPC and CSS, the stress evolution in the periphery rock mass of the retreat channel at the final mining is examined, and the influence of these two factors on the stress evolution law of the surrounding rock of the retracement channel under the goaf is obtained.

1. With different LPC as the research index, in the process of final mining, the surrounding rock stress of the withdrawal channel under the goaf has the following evolution law with the change of the LPC: with the increase of the LPC, the PAP of the S-SR of the withdrawal channel increases and stabilizes in a certain range; the variation trend of the principal stresses of the AWF under different LPC is the same, showing a trend of increasing first, then decreasing and then increasing, which is similar to the ' N ' shape ; with the increase of the LPC, the MDS of the S-SR of the withdrawal channel decreases first and then increases, and stabilizes in a certain range.

2. Taking different CSS ( no overlying goaf, with a 20 m separation from the overlying goaf and with a 75 m separation from the overlying goaf ) as the research index, in the process of final mining, the surrounding rock stress of the withdrawal channel under the goaf has the following evolution law with the change of CSS: with the increase of the distance between the withdrawal channel and the overlying goaf, the PAP of the S-SR of the withdrawal channel decreases first and then increases ; the change trend of the principal stresses of the AWF without overlying goaf and with a 75 m separation from the overlying goaf increases first, then decreases, and remains within a certain range, and the change trend of the principal stresses of the AWF 20 m away from the overlying goaf increases first and then decreases. With the increase of the CSS, the MDS of the S-SR of the withdrawal channel decreases first and then increases.

From the above research results, it can be seen that the influence of the LPC on the surrounding rock stress of the retracement channel under the goaf gradually decreases with the increase of the LPC; with the increase of CSS, the stress state of the surrounding rock of the withdrawal channel under the goaf is gradually similar to that of the non-overlying goaf.

Similar test of the withdrawal channel at the final mining stage under the goaf

The stability of the surrounding rock in the retreat mining channel under the goaf at the final mining stage is investigated via a similarity simulation experiment, with the objective of facilitating the safe and highly efficient withdrawal of equipment.

According to the size of the experimental area and the test bench, the Model similarity ratio is 1:100. The test uses sand, lime and gypsum as the main proportioning material, and mica sheet as the layered material.

As the final mining stage closes in on the withdrawal channel (from 30 m to 0 m), the deformation of the withdrawal channel under the four stages of coal pillar width of 30 m, 20 m, 10 m and 0 m is analyzed respectively, as shown in Fig. 13.

Fig. 13.

Similar test of withdrawal channel in final mining stage under goaf

Research findings indicate that the stability of the surrounding rock in the retreat mining channel under the goaf is governed by the MDSZ. This zone manifests as the primary deformation area in similarity simulation experiments, while appearing as the stress concentration area in numerical simulations. Upon the connection between the retreat channel and the mining face, the overlying strata undergo rapid subsidence, triggering an abrupt stress change and severe roadway deformation. Comparative experimental results demonstrate that pre-splitting and roof cutting technology can effectively mitigate this issue, serving as a critical technique for ensuring safe equipment retreat.

Engineering test

Based on the on-site investigation results, at the final mining of the 5106 working face in the No. 5 coal seam at this mine, the original support scheme failed, leading to significant deformation of the withdrawal roadway. It not only severely threatens equipment withdrawal but greatly slows down the relocation of the working face, impacting coal mine production. As a result, a section of the withdrawal roadway, about 50 m long, was chosen as the on-site engineering test section.

Control mechanism

Since the MDSZ can captures the progression of rock mass from elastic to plastic deformation, in studying the evolution of the deviatoric stress of the surrounding rock of the withdrawal roadway under the goaf, it is found that at the final mining, the MDSZ in the surrounding rock of the withdrawal roadway is in the transition zone between the roof and the S-SR. This area should be prioritized for control. To prevent the out-of-control cutting along the coal wall in this region, a cable anchor reinforcement support method in the peak area of deviatoric stress, along with active control of the roof fracture position, is proposed. The high - prestress field formed by the cable anchor in the surrounding rock, on the one hand, transforms the tensile stress in the surrounding rock into compressive stress, improving the stress state of the surrounding rock; on the other hand, through the shear - resistance performance of the cable anchor itself and the enhancement of the normal binding force of the structural plane of the surrounding rock, the shear - resistance ability of the surrounding rock can be significantly improved. Therefore, cable anchors can be used to support the MDSZ of the surrounding rock of the withdrawal roadway, so that the support range covers the MDSZ and is anchored to the elastic zone of the surrounding rock. In addition, when the fracture line of the rock block is located behind the hydraulic support, it is most beneficial for the control of the surrounding rock of the withdrawal roadway. It can shorten the overhanging distance of the roof, reduce the load borne by the surrounding rock of the withdrawal roadway, and is also the safest for the personnel and equipment in the withdrawal roadway. A pre-splitting roof cutting scheme is adopted for the roof on the G-SR. As the longwall face is a considerable distance from the retreat gateway, the roof of the G-SR is blasted to generate cracks in the roof at the predetermined position in advance, so that the roof can break smoothly after the hydraulic support passes through the predetermined position.

Supporting scheme

The original support scheme for the withdrawal channel employed symmetric support, which was somewhat reasonable. However, it failed to account for the support needs in the MDSZ. By studying the support mechanism for this peak area of the withdrawal channel, it was determined that the critical support zone was the peak of the surrounding rock’s deviatoric stress. Since this area is 6.65 to 9.42 m from the withdrawal channel, an asymmetric strengthening support was chosen, using anchor cables of appropriate length and strong performance.

In the original support scheme, roof support was provided primarily by rock bolts, wire mesh, and cables. The roof used Φ20 × 2000 mm deformed steel bolts with a layout density of 5 bolts a row, a spacing 800 × 800 mm. The bolts on both sides extended toward the sidewall at an angle of 15°, and they were fitted with 150 × 150 × 10 mm pallets. The specification of the cables was Φ17.8 × 7200 mm, with a layout density of 3 cables a row and spacing of 1500 × 2000 mm, and a 30 × 30 mm diamond-shaped metal mesh was used for surface protection. For the sidewall support, Φ20 × 2000 mm deformed steel bolts were chosen, arranged with three bolts a row and spaced 800 × 800 mm apart. The end bolts were uesd at an angle of 15°, and the pallet specification was 150 × 150 × 10 mm.Through the analysis of the key support mechanism for the MDSZ in the withdrawal channel under the goaf, it was found that the MDSZ in the withdrawal channel was the intersection area between the roof and the non - mining sidewall, and reinforcement support was required for this area. The reinforcement anchor cable on the roof was 300 mm away from the non-mining sidewall, and the angle with the horizontal direction was 55°. The specification of the reinforcement anchor cable was Φ21.6 mm×9200 mm; The reinforcement anchor cable on the non-mining side was 400 mm away from the roof, and the angle with the horizontal direction was 25°. The specification of the reinforcement anchor cable was Φ21.6 × 10,500 mm, and the pallet specification was 300 × 300 × 16 mm, as shown in Fig. 14.

According to the original support scheme and the strengthened support scheme, FLAC3D prestress numerical simulations were carried out respectively, and the prestress fields formed by the bolts and cables were compared. In initial scheme, the prestress field only formed a strengthened structure in the nearby area. The 0.01 MPa pressure arch formed was 6.4–6.9 m away from the farthest side of the MDSZ, and the 0.02 MPa pressure arch was even farther from the MDSZ; After the reinforcement support of the MDSZ in the withdrawal channel with reinforcement anchor cables, the strengthened area formed by the prestress field on the non - mining sidewall of the withdrawal channel increased significantly. The 0.01 MPa pressure arch formed covered more than 75% of the MDSZ, and the maximum distance from the farthest side of the MDSZ was 1.7 m. The 0.02 MPa pressure arch could also reach the MDSZ. In the strengthened support scheme, the areas of the 0.01 MPa and 0.02 MPa pressure arches were significantly larger than those in the original support scheme. Moreover, the 0.01 MPa pressure arch had covered most of the MDSZ, which enhanced the shear resistance of the weak areas of the surrounding rock support.

The bidirectional energy-concentrating blasting technology uses a bidirectional energy-concentrating tube to directionally release the high-pressure gas generated by blasting, thereby generating directional fractures in rock mass. Finally, continuous fractures are generated in the direction of the axis of the energy-concentrating hole to achieve directional pre-splitting. In order to make the roof break behind the hydraulic support, as the longwall face advanced to within 200 m of the retreat roadway, a blasting hole L1 was drilled into the roof. The L1 was angled at 51° to the horizontal, the hole depth was 17.2 m, the diameter was 48 mm, the charging length was 12 m, and the hole-sealing length was 5.2 m. After blasting, fractures were formed in the roof rock stratum. After the hydraulic support passed through this area, the roof broke at the predetermined position. A gangue-blocking metal mesh was laid behind the hydraulic support to prevent gangue from entering the working face.

Fig. 14.

Comparison of the surrounding rock reinforcement areas of the combined bolt and cable support between the original support and the partitioned asymmetric support scheme for key regions.

Mine pressure monitoring

The effect of the basic support of bolts and cables in the withdrawal roadway under the goaf, the strengthened support in the MDSZ, and the pre-splitting roof cutting on the G-SR is shown in Fig. 14. After the construction was completed, regular effect observation and deformation monitoring were carried out. It was found that the surrounding rock had good integrity, and there were no obvious phenomena of surrounding rock fragmentation and large deformation. The bolts and anchor cables functioned well, and there were no phenomena of failure, anchor disengagement, and anchor falling. Measuring instruments such as laser rangefinders and leveling staffs were adopted. Considering the actual conditions of the underground roadway, equipLment with high accuracy and convenient data collection was selected. The displacement of the non-mining side and the roof was observed using the cross-point layout method. During the final mining, the sidewall displacement was between 100 mm and 150 mm, and the roof displacement was about 150–200 mm, as shown in Fig. 14. During the withdrawal period, the deformation of the withdrawal roadway was stable, and the equipment was withdrawn smoothly.

Conclusion

1. Variation principle of surrounding rock stress in the retreat roadway during the final mining under the goaf under different LPC and CSS.

1. At the final mining period, variation principle of surrounding rock stress under the goaf with different LPC follows specific patterns: as the LPC increases, the PAP on the S-SR rises and stabilizes within a certain range. The variation trends of the principal stresses ahead of the AWF with different LPC are similar, showing a pattern of first increasing, then decreasing, and then increasing again, roughly forming an “N” shape.

2. At the final mining period, variation principle of surrounding rock stress under the goaf with different CSS follows these patterns: as the distance from the overlying goaf increases, the PAP on the non-mining side of the withdrawal roadway first decreases and then increases; the trends of the maximum and minimum principal stresses ahead of the AWF without an overlying goaf and at 75 m from the goaf initially increase, then decrease, and stay within a certain range. Meanwhile, the stresses at 25 m from the goaf first increase, then decrease, then increase again. Additionally, as the CSS increases, the MDS on the non-mining side of the withdrawal roadway first decreases and then increases.

3. The influence of the LPC on the surrounding rock stress of the withdrawal roadway under the goaf gradually diminishes as the LPC increases. As the CSS widens, the stress state of the surrounding rock in the withdrawal roadway under the goaf gradually becomes similar to that without an overlying goaf.

• (2)The MDSZ in the surrounding rock within the withdrawal channel is concentrated at the intersection of the roof and the non-mining side. A targeted scheme for “anchor cable reinforcement support in the MDSZ” is proposed. In comparison with the initial support design, the pressure arch areas of 0.01 MPa and 0.02 MPa in the strengthened support for the withdrawal channel are significantly increased, with the 0.01 MPa pressure arch covering most of the MDSZ. This enhances the shear resistance of the weak support areas of the surrounding rock. Additionally, pre-splitting roof cutting is used on the G-SR, which shortens the roof overhang and reduces the load on the surrounding rock of the withdrawal channel. Ultimately, the displacement of the side wall ranges between 100 mm and 150 mm, while the roof displacement is about 200 mm, ensuring the safe withdrawal of equipment.

Author contributions

Conceptualization: D.C. and W.H.Formal analysis: D.C., J.T. and C. G.Funding acquisition: D.C. and W.H.Investigation: D.C. and W.H.Methodology: D.C. and Z.Z.Resources: D.C., J.T. and W.H.Software: D.C., J.T., C. G. and Z.Z.Supervision: D.C. and Z.Z. Validation: D.C. and W.H.Visualization: D.C. Writing – original draft: D.C., J.T. and Z.Z.Writing – review & editing: D.C., J.T. and C. G.

Funding

National Natural Science Foundation of China (52374149, 52004286, 51974317), Jiaxing Science and Technology Planning Project (2024AY40014), General Scientific Research Project of the Education Department of Zhejiang Province (Y202352229).

Data availability

All data generated or analysed during this study are included in this published article.

Declarations

Competing interests

The authors declare no competing interests.