Characteristics of Mine Pressure Behavior and Zoned Support Technology for Advancing Working Face in Ultra-Close Coal Seams

Abstract

To address the issues of severe surrounding-rock failure and ground support component failure in advancing working-face driving roadways (AWFDRs) in ultra-close coal seams, this study used the 5202 air-return roadway in Huaye Coal Mine as a case study and for engineering background. Numerical simulation, theoretical analysis, and industrial application methods were adopted to analyze the laws of the dynamic evolution of vertical stress in such roadways. The mine pressure behaviors of AWFDRs in ultra-close coal seams were also clarified, thereby enabling the proposal of a solution; namely, zoned support technology. The results show that the 5202 air-return roadway, as an AWFDR in an ultra-close coal seam, exhibits five different characteristic behaviors of mine pressure zones during excavation. Zone 1 is influenced by the adjacent working-face mining under goaf; Zone 2 is influenced by the adjacent goaf lateral abutment stress under goaf; Zone 3 is influenced by the stress of the overlying solid coal; Zone 4 is influenced by the adjacent goaf lateral abutment stress under the overlying solid coal; and Zone 5 is influenced by stabilized stress under the overlying solid coal. The mine pressure behaviors of these zones were ranked, from most intense to weakest, as follows: Zone 3 > Zone 1 > Zone 4 > Zone 2 > Zone 5. Based on this, a basic support scheme was proposed, which involves using bolt–mesh–beam supports combined with shed supports under the goaf and bolt–mesh–beam supports combined with roof anchor cables under the overlying solid coal. Additionally, in Zones 1 and 3, roof anchor cables or rib anchor cables were supplemented as reinforcing supports, which were combined with the basic support scheme described above to form a zoned support scheme for the AWFDR. The analysis of mine pressure behavior and implementation of a zoned support scheme for AWFDRs in ultra-close coal seams provides technical and engineering references for roadway supports under similar mining conditions.

1. Introduction

Energy is the source of power for modern economic growth and forms a crucial material foundation for fostering sustainable development within national economies. Recently, global energy production and consumption have been increasing year on year [1,2]. As a vital resource for energy production, coal also saw its production and consumption reach an all-time high in 2023 [3]. However, in reality, coal mining is concentrated in countries such as China, India, Indonesia, and Australia. Among these, China is the world’s largest coal consumer, accounting for 56% of global coal consumption.

As China’s coal consumption increases year on year, the intensity of coal mining also increases day by day [4,5]. This continuous increase in production will lead to the tension of mining alternation, especially in single-wing mining areas, as both the upper section of the working face and the lower section of the required roadway layout are mined simultaneously; the area adjacent to the upper section of the working face of the roadway is denoted as the “AWFDR” [6,7,8]. Due to the fact that the mining roadway is influenced by the adjacent working face, the deformation of the rock surrounding the roadway is substantial, and adequately supporting this rock is difficult [9,10,11,12]. Many scholars have carried out a large number of research studies—ranging from theory-based research to numerical simulation and on-site experiments—focused on controlling the rock surrounding the roadway when mining and digging. Kang Hongpu et al. [13,14,15] revealed the spatial and temporal distribution of stress peaks in the AWFDR through field measurement and a comparative analysis of multiple mines; the stress peak usually occurs behind the goaf. Chen Dingchao et al. [16] and Wang Meng et al. [17] studied the headway under wide and narrow coal pillar conditions, determined a reasonable coal pillar size, and put forward the principle of dynamic segmentation control and technical solutions; Liu Jinhai et al. [18] proposed a research approach for AWFDRs in deep and thick coal seams, which involved not only determining whether the advanced abutment pressure ahead of the working face overlaps with the lateral abutment pressure outside the goaf and the degree of such a superposition, but also involved identifying the location of high-stress zones formed and optimizing the support scheme used based on these findings; Wang Yongtao et al. [19] analyzed the deformation mechanisms of gob-side roadways using theoretical models and numerical simulations, identifying critical block fractures at 14.07 m and stress reduction zones within 16.12 m. They proposed an optimized support scheme combining high-strength bolts/cables with mesh/steel belts to control asymmetric deformation; this was validated using field monitoring, showing reduced displacements of 61~154 mm. Dai Lianpeng et al. [20] found that geostress exceeding a critical value is a key driver for the formation of rockbursts and that rock brittleness, uniaxial compressive strength, and the amount of roadway excavation undertaken significantly influence the convergence of surrounding rock and the occurrence of rockbursts. They also proposed that the design criterion for support in rockburst-prone roadways is determined by the intersection of the surrounding rock convergence curve and the support squeezing deformation curve. Wang Yuliang et al. [21] studied the failure and movement characteristics of overlying strata under thick aquifers. These findings are relevant to mine pressure behavior in roadways driven towards the working face: the periodic breaking of overlying strata induces dynamic loads, water pressure softens surrounding rocks, and fracture development affects roadway stability.

The above research has mainly focused on AWFDRs in single-seam mining, and there are fewer research results related to close coal seam groups. In fact, most of the coal seams in mining areas are in the form of coal seam groups [22,23,24] and, after mining of the upper seams occurs, the lower seams are affected; in particular, during the mining of the lower seams of ultra-close coal seam groups, these lower seams are affected by the damage caused by the mining of the upper seams [25,26] and by the loading of the coal pillars [27,28,29], meaning that control of the rock surrounding AWFDRs is more difficult. Studies investigating this particular condition are relatively rare. Based on this, using the Huaye Coal Industry’s AWFDR in an ultra-close coal seam as the engineering background, numerical simulation and theoretical analysis were used to analyze the dynamic evolution of vertical stress in the rock surrounding an AWFDR in an ultra-close coal seam. This study also clarified the characteristics of mine pressure behavior in this type of roadway, proposed a segmental support technology scheme for the 5202 air-return roadway under complex mining conditions, and carried out an industrial test in order to provide technical and engineering support solutions for roadway supports in similar projects under similar engineering conditions.

2. Project Overview

Huaye Coal Mine, Lvliang City, Shanxi Province, China mines the 4^# and 5^# coal seams, with the distance between the No. 4 and No. 5 coals seams being approximately 2.2 m, characterizing them as belonging to an ultra-close coal seam group. The average thickness of the 4^# coal seam is 2.23 m, the average thickness of the 5^# coal seam is 3.48 m, the dip angle is 2–8°, and the average depth of the 5^# coal seam is 305 m. The distribution of the top and bottom rock layers of the 4^# and 5^# coal seams and their physical and mechanical parameters are shown in

Table 1. Distribution and physical–mechanical parameters of roof and floor strata in No. 4 and No. 5 coal seams.

Lithology	Layer Thickness/m	Densities/kg·m−3	Compressive Strength/MPa	Tensile Strength/MPa	Cohesion/MPa	Angle of Internal Friction ϕ/°	Modulus of Elasticity/GPa	Poisson’s Ratio
Medium-grained sandstone	4.48	2593	36.6	2.54	9.00	33.66	12.0	0.22
Siltstone	5.52	2588	36.8	1.71	8.70	36.56	15.4	0.20
Mudstone	3.80	2582	27.3	0.61	4.30	36.69	7.7	0.23
Sandy mudstone	5.90	2613	33.6	1.46	4.10	36.51	10.0	0.21
4# coal	2.23	1425	8.2	0.38	1.18	21.81	1.7	0.22
Mudstone	2.20	2592	22.4	0.76	4.30	35.35	8.0	0.26
5# coal	3.48	1393	8.8	0.38	1.28	25.10	1.6	0.23
Sandy mudstone	2.26	2652	30.5	1.22	6.30	30.93	11.5	0.24
Medium-grained sandstone	8.38	2598	52.1	1.90	6.40	33.57	11.3	0.20
Mudstone	4.39	2580	26.8	0.85	4.80	35.54	8.2	0.21
Limestone	7.23	2622	57.6	1.89	7.70	36.46	14.3	0.22

.

At present, the working face of the 5^# seam, 5201, is being mined back; the strike length of the working face is 1306 m and the inclination length is 158 m. The eastern boundary of the working face of 5201 is the protective coal pillar at the boundary of the well field, the western boundary is the three main mining entry panels, the southern boundary is the solid coal, and the northern boundary is the working face of 5202 in the 5^# seam (the successor of the working face of 5201). A total of 25 m of protective coal pillar is set aside for the section in the middle of the working face, and the engineering plan for its mining is shown in Figure 1.

Among them, the upper part of the 5202 air return roadway (mileage: 0~990 m (J₁-J₃)) comprises the 4201 goaf of the 4^# coal seam; the section at 990~1663 m (J₃-J₄) is the solid coal area of the 4^# coal seam. According to the plans, it is necessary to construct the 5202 working face roadway during the mining of the 5201 working face. For this reason, the 5202 air return roadway will experience four mining conditions during the excavation process: ① J₁-J₂: as shown in Figure 2a, this is driven under the 4201 goaf along the solid coal seam of the 5201 working face; ② J₂-J₃: as shown in Figure 2b, this is driven under the 4201 goaf along the goaf of the 5201 working face; ③ J₃-J₄: as shown in Figure 2c, this is driven under the solid coal of seam 4 along the goaf of the 5201 working face; ④ J₄-J₅: as shown in Figure 2d, this is driven under the solid coal of seam 4 along the solid coal of Seam 5. Among these, J₁ is the opening position of the roadway, J₂ is the convergence position of the mining and digging working faces, J₃ is the open–off cut of the 4201 working face of the 4^# coal seam, J₄ is the open–off cut of the 5201 working face, and J₅ is the open–off cut of the 5202 working face.

3. Numerical Simulation

3.1. Model Building

In order to study the stability of the surrounding rock, under different conditions, in the 5202 air return roadway (Huaye Coal Mine), a numerical simulation study was carried out using FLAC^3D 6.0 software. Based on the actual engineering conditions of the 4201, 5201, and 5202 working faces, the numerical simulation model was constructed as shown in Figure 3. The size of the established model is 500 m (strike) × 150 m (inclination) × 60 m (layer thickness). In order to simulate the stability of the roadway at different stages, the simulation length of the 5202 air-return roadway is 440 m, and the roadway section is 4.5 m (width) × 2.9 m (height) along the base plate; the length of the 5201 working face is 420 m. In the numerical model, the stress and displacement boundary conditions are set as follows: ① Stress boundary condition—A uniform vertical load of 6.5 MPa is applied to the top surface of the model to simulate the overburden self-weight stress, which directly represents the vertical in situ stress component considered in the model. This load magnitude is derived from calculations based on the formation’s burial depth and rock unit weight in the study area, defining the initial vertical in situ stress state of the target layer. ② Displacement boundary conditions—The bottom boundary is set as a fully fixed constraint (both the normal and tangential displacements are zero), while the lateral boundaries are constrained to prevent horizontal normal displacements (normal displacement = 0) with free tangential displacements, mimicking the realistic boundary constraints of the geological formation.

The simulation process is as follows: initial equilibrium → the 4201 working face is mined back and the equilibrium is calculated → the 5201 working face is mined back for 270 m → the 5202 air-return roadway and the 5201 working face are advanced in opposite directions simultaneously. The ratio of the digging speed of the 5202 air-return roadway to the mining-back speed of the 5201 working face is 2:1; → the equilibrium is calculated.

3.2. Vertical Stress Analysis

The vertical stress field distribution of the rock surrounding the mining and digging working faces is shown at different times in Figure 4. It can be seen that the 5202 air-return roadway is influenced by the mining of the adjacent 5201 working face and the adjacent goaf lateral abutment stress. Additionally, the stress of the roadway-side goaf is clearly higher than that of the side of solid coal, and the stress under the solid coal of the 4^# coal seam is higher than that under the goaf, especially near the junction of the two, where the stress concentration is the most obvious. We then analyzed the evolution patterns of vertical stress in the rock surrounding the roadway in two sections: before and after the convergence of the mining and excavation faces, and before and after entering the solid coal of Seam 4.

3.2.1. Vertical Stress Evolution Pattern Before and After Face Convergence

The stress measurement line is arranged in the tendency direction at the convergence position of the mining face (point J₂, y = 150), and the vertical stress curves of both rib sides and the coal pillar at different distances from the working face are shown in Figure 5. With the mining and digging faces gradually approaching convergence, the vertical stress in the rock surrounding the roadway shows different distribution characteristics, as follows: ① Beyond 40 m ahead of the advancing face, the stress of the rib sides of both the roadway and the coal pillar grows more slowly; ② 40 m behind the advancing longwall face, influenced by the advanced abutment pressure from the working face, the stress in both rib sides and in the coal pillar increases rapidly; ③ within the distance from the face convergence point to 20 m behind the working face, the rate of the increase in coal column stress increases further, and reaches its maximum value at 20 m behind the working face. The peak stress in the coal pillar is 26.8 MPa on the side adjacent to the 5201 goaf and 19.7 MPa on the side near the 5202 air-return roadway, representing increases of 288% and 252%, respectively, compared to the 40 m advanced position. Additionally, the stress levels of both rib sides of the 5202 air-return roadway exhibit asymmetric distribution: the stress on the coal pillar side is 2.1 MPa higher than that on the solid coal side. At this stage, the 5202 air-return roadway is subjected to the most intense mining-induced effects from the adjacent 5201 working face; ④ within the range of 20~80 m behind the 5201 working face, the mining-induced effects gradually weaken, and the stress in the surrounding rock of the roadway decreases progressively; ⑤ 80 m behind the working face, the stress curve flattens out. Influenced by the residual abutment pressure from the adjacent goaf, the peak vertical stresses on the coal pillar stabilize at 15.8 MPa and 10.7 MPa, respectively.

3.2.2. Vertical Stress Evolution Patterns Before and After Entering the Solid Coal Section of the 4^# Coal Seam

In order to study the evolution patterns of vertical stress before and after the roadway enters the solid coal section of the 4^# seam, two peak stress levels of the coal column in the section, each along the direction of the stress measurement line (y = 310~470), as well as the vertical stress curve of the coal column, were measured, as shown in Figure 6. The results indicate that, before J₃, a stress-reduced zone exists in the floor of the goaf in Seam 4, with coal pillar stress decreasing slowly at this point. After J₃, the coal pillar stress increases rapidly, reaching its peak at 9 m behind J₃, followed by a decline. The peak stress levels on the goaf side and the air-return roadway side are 32.2 MPa and 24.1 MPa, respectively, representing increases of over 20% compared to the peak stress during dynamic pressure effects from the 5201 working face. At 25 m behind J₃, the stress stabilizes but remains slightly higher than before J₃. After J₄, a stress peak only forms near the roadway side, which is decreased compared to pre-J₄ conditions.

4. Mechanism of the Effect of Overlying Solid Coal on Subgrade Stresses

In the above numerical simulation study, it was found that the peak stress of the rock surrounding the 5202 air-return roadway of the lower coal seam occurs near point J₃; in order to further study the influence of the overlying solid coal on the stress of the rock surrounding the roadway of the lower coal seam, a mechanical analysis was carried out on the coal rock body near point J₃. According to the semi-planar body theory [30,31,32,33], the mathematical model of the additional force on any point P(x,y) of the solid coal bottom plate is established using the lower end point of the solid coal boundary as the coordinate origin, as shown in Figure 7.

In Figure 7, γ and H denote the average bulk weight and the average depth of the overlying rock layer in kN/m³ and m, respectively; K represents the stress concentration coefficient of solid coal; and L₁~L₃ are the lengths of the unloading protection zone of the goaf, the plastic zone of the solid coal, and the elastic pressurized zone of the solid coal, respectively, in m.

According to the stress characteristics of the air-mining zone and solid coal in Figure 7, the additional load calculation formulae for each section are shown below.

The additional load of the pressure relief protection zone of the air-mining zone is as follows:

$$q_{L_1}=-{\displaystyle \frac{\gamma H}{L_1}}\xi -\gamma H$$

(1)

The additional loads in the plastic zone are calculated as follows:

$$q_{L_2}={\displaystyle \frac{K\gamma H}{L_2}}\xi +\gamma H$$

(2)

The additional loads in the elastic pressurized zone are calculated as follows:

$$q_{L_3}=-{\displaystyle \frac{\left(K-1\right)\gamma H}{L_3}}\xi +{\displaystyle \frac{\left(K-1\right)\gamma H(L_2+L_3)}{L_3}}$$

(3)

When taking a tiny length, dξ, on the boundary of the semi-plane body and considering the additional force, dF = qdξ, applied on it as a tiny concentrated force, the additional stress induced at the point P(x,y) is as follows:

$$\left\{\begin{array}{c}d{{\sigma }^{\prime }}_x=-{\displaystyle \frac{2q\mathrm{d}\xi }{\pi }}{\displaystyle \frac{y^3}{{\left[y^2+{(x-\xi )}^2\right]}^2}}\\ d{{\sigma }^{\prime }}_y=-{\displaystyle \frac{2q\mathrm{d}\xi }{\pi }}{\displaystyle \frac{y{(x-\xi )}^2}{{\left[y^2+{(x-\xi )}^2\right]}^2}}\\ d{{\tau }^{\prime }}_{xy}=-{\displaystyle \frac{2q\mathrm{d}\xi }{\pi }}{\displaystyle \frac{y^2(x-\xi )}{{\left[y^2+{(x-\xi )}^2\right]}^2}}\end{array}\right.$$

(4)

where ${{\sigma }^{\prime }}_x$ is the horizontal additional stress in MPa; ${{\sigma }^{\prime }}_y$ is the vertical additional stress in MPa; ${{\tau }^{\prime }}_{xy}$ is the additional shear stress in MPa; and x and y are the vertical and horizontal distances from the concentrated force F to the point P, respectively.

According to the principle of stress superposition, the additional force at any point P of the bottom plate is as follows:

$$\left\{\begin{array}{c}{{\sigma }^{\prime }}_x=-{\displaystyle \frac{2y^3}{\mathsf{\pi }}}\left\{{\displaystyle {\int }_{L_2}^{L_2+L_3}\frac{q_{L_3}\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}+{\displaystyle {\int }_0^{L_2}\frac{q_{L_2}\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}+}{\displaystyle {\int }_{-L_1}^0\frac{q_{L_1}\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}}}\right\}\\ {{\sigma }^{\prime }}_y=-{\displaystyle \frac{2y}{\mathsf{\pi }}}\left\{{\displaystyle {\int }_{L_2}^{L_2+L_3}\frac{q_{L_3}{\left(x-\xi \right)}^2\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}+{\displaystyle {\int }_0^{L_2}\frac{q_{L_2}{\left(x-\xi \right)}^2\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}+}{\displaystyle {\int }_{-L_1}^0\frac{q_{L_1}{\left(x-\xi \right)}^2\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}}}\right\}\\ {{\tau }^{\prime }}_{xy}=-{\displaystyle \frac{2y^2}{\mathsf{\pi }}}\left\{{\displaystyle {\int }_{L_2}^{L_2+L_3}\frac{q_{L_3}\left(x-\xi \right)\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}+{\displaystyle {\int }_0^{L_2}\frac{q_{L_2}\left(x-\xi \right)\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}+}{\displaystyle {\int }_{-L_1}^0\frac{q_{L_1}\left(x-\xi \right)\mathrm{d}\xi }{{\left[y^2+{\left(x-\xi \right)}^2\right]}^2}}}\right\}\end{array}\right.$$

(5)

After adding the additional stress at point P to the original rock stress, the force at any point of the solid coal footing can be obtained as follows:

$$\left\{\begin{array}{c}{\sigma }_x={\sigma }_x^{\prime }+\gamma H\\ {\sigma }_y={\sigma }_y^{\prime }+\lambda \gamma H\\ {\tau }_{xy}={\tau }_{xy}^{\prime }\end{array}\right.$$

(6)

where ${\sigma }_x$ denotes horizontal stress in MPa; ${\sigma }_y$ denotes vertical stress in MPa; and ${\tau }_{xy}$ represents shear stress in MPa.

When combined with the engineering and geological conditions of Huaye Coal Industry, γ is set as 25 kN/m³, H is set as 300 m, K is set as 3.4, and the mean value is set as 63 m, according to L₁ = (0.12~0.3) H [34].

According to the limit equilibrium theory [35,36], the length of the plastic zone of the solid coal, L₂, and the length of the elastic pressurized zone, L₃, are calculated as follows:

$$L_2={\displaystyle \frac{\lambda M}{2\tan \varphi }}\ln \left({\displaystyle \frac{K\gamma H+\frac{C}{\tan \varphi }}{\frac{C}{\tan \varphi }+\frac{P_0}{\lambda }}}\right)$$

(7)

where λ denotes the lateral pressure coefficient and is set as 0.28; M represents the mining height, set as the coal thickness of 2.23 m; φ denotes the friction angle within the coal body, set as 21.8°; C denotes the cohesion of the coal body, set as 1.18 MPa; and P₀ represents the strength of the coal-side support, set as 0.1 MPa.

$$L_3={\displaystyle \frac{\lambda M}{f}}\ln {\lambda }_1{\displaystyle \frac{1+\sin \varphi }{1-\sin \varphi }}$$

(8)

where f is the interlayer friction coefficient, which is set as 0.26.

Each parameter is substituted into Equations (7) and (8) to obtain L₂ = 3.08 m and L₃ = 11.43 m.

Each parameter is substituted into Equations (7) and (8) and calculated using numerical calculation software to obtain the stress distribution maps at different depths of the solid coal bottom plate, as shown in Figure 8. The results show that vertical, horizontal, and shear stresses will be concentrated under the solid; for example, in the area where the roadway is located, namely in the top plate, the peak vertical stress is about 12 MPa, and the peak stress is about 10 m away from the edge of the solid coal (point J₃). Additionally, as we move away from the edge of the solid coal, the degree of concentration of the stress under the solid coal decreases gradually; and it is basically unaffected by the solid coal and restores to the original rock stress levels after 20 m. The stress concentration under the solid coal gradually decreases as we move away from the edge of the solid coal. In addition, the vertical stress is about 5 MPa on the side of the goaf adjacent to point J₃, forming a stress reduction zone.

The results of the theoretical analysis and the numerical simulation are basically consistent. Before the roadway enters the solid coal section of the 4^# seam, the surrounding rock stress decreases; the stress rises rapidly to a peak and then decreases after entering the solid coal section; finally, it tends back to the original rock stress levels. This shows that the stress concentration of the overlying solid coal on the lower coal seam’s 5202 air-return roadway is mainly in the range of 0~20 m, that the influenced range is related to the length of the plastic and elastic pressurized zones on the side of the overlying solid coal section, and that the influence range is about 1.4 times the sum of the length of both of these zones. In this area, the pressure of the rock surrounding the roadway is large, so it is necessary to strengthen the supports for the top plate and the two sides within this range to ensure the stability of the roadway’s surrounding rock.

5. Analysis of the Characteristic Mine Pressure Behaviors of AWFDR in an Ultra-Close Coal Seam

Through the above analysis, due to different mining conditions in the surrounding rock, the 5202 air-return roadway, also known as the AWFDR in ultra-close coal seams, was shown to exhibit five different characteristic behaviors observed in mine pressure zones during excavation. The details are as follows:

(1)

Zone 1: influenced by the adjacent working face under the goaf.

Zone 1 is within 80 m from the opening of the roadway to the adjacent working face. The roadway is influenced by the mining of the adjacent working face in this zone; specifically, after the roadway enters the influence range of the advanced abutment pressure ahead of the working face, the pressure of the rock surrounding the roadway gradually increases, and the amount of deformation increases. After the convergence of the mining and digging working faces, the roadway is driven under the goaf. Simultaneously, along the goaf, the overlying strata structure in the goaf remains unstable, with the main roof breaking and rotating while subsiding, forming lateral abutment pressure on the goaf-adjacent side of the coal pillar and triggering severe deformation and failure of the roadway area [37]. Additionally, at this time, the surrounding rock deformation rate significantly increases, necessitating the reinforcement of the support on the coal pillar side of the roadway in this zone.

(2)

Zone 2: influenced by the adjacent goaf lateral abutment stress under the goaf.

Zone 2 spans from 80 m behind the working face to the edge of the overlying goaf. The roadway is driven along the adjacent goaf, under the goaf. As it moves away from the rear of the working face, this zone is no longer affected by mining-induced influences from the adjacent working face. The collapsed waste rock in the adjacent goaf gradually becomes compacted by the overlying strata, and, due to the unloading effect of the overlying goaf, the surrounding rock pressure and deformation decrease compared to the previous zone.

(3)

Zone 3: influenced by the stress concentration of the overlying solid coal.

Zone 3 refers to the range, within 0–20 m, wherein the roadway enters the overlying solid coal section. Under the influence of concentrated loads from the solid coal outside the overlying goaf [35], the surrounding rock stress in the roadway increases significantly, leading to a steep rise in mine pressure behaviors.

(4)

Zone 4: influenced by the adjacent goaf lateral abutment stress under the overlying solid coal.

Zone 4 spans the range from 20 m into the solid coal to the adjacent working face’s starting cut, with the roadway driving along the goaf under the solid coal. As it moves away from the edge of the overlying solid coal, the stress field in the overlying solid coal tends to return to its original stress state. This zone is similar to traditional goaf-side entry drivage, with more mine pressure behaviors demonstrated than in zone 2.

(5)

Zone 5: influenced by stabilized stress under the overlying solid coal.

Zone 5 is the range wherein, after passing the adjacent working face’s starting cut, the roadway is driven along the solid coal and under the solid coal, representing conventional roadway drivage. Mine pressure behaviors in this zone are minimal compared to in the previous four zones.

Based on the above analysis, it can be observed that the drivage of AWFDRs in ultra-close coal seams is mainly influenced by the adjacent goaf lateral abutment stress and by concentrated loads from the solid coal outside the overlying goaf. The mine pressure behaviors of these zones rank, from most intense to weakest, as follows: Zone 3 > Zone 1 > Zone 4 > Zone 2 > Zone 5.

6. Roadway Zone Support Scheme

Based on the above analysis, on-site engineering practice experience, and the actual conditions of the roadway, the 5202 air-return roadway was divided into five zones, taking into account the characteristics of mine pressure behavior as zoning support criteria, as shown in Figure 9. Due to the small distance between the two coal seams, it was unsuitable to use anchor cable supports for the roof of the roadway under the goaf; therefore, a basic support scheme was proposed, which involves using bolt–mesh–beam supports combined with shed supports under the goaf, and bolt–mesh–beam supports combined with roof anchor cables under the overlying solid coal. Additionally, in Zone 1 and 3, roof anchor cables or rib anchor cables were supplemented as reinforcing supports, which were combined with the basic support scheme described above to form a zoned support scheme for the AWFDR. The specific support scheme is as follows.

(1)

Zone 1 (opening of the roadway—40 m before face convergence point):

The roof and both ribs are supported by bolts with a specification of φ20 mm × 2000 mm, installed at a spacing of 800 mm × 900 mm (transverse × longitudinal). The steel I-beam supports are arranged at 0.9 m intervals.

The coal pillar side is reinforced with short anchor cables (φ17.8 mm × 4300 mm), installed in pairs per row at a spacing of 1600 mm × 1800 mm (transverse × longitudinal), as shown in Figure 10a.

The roadway zone will be influenced by the adjacent 5201 working face; at the same time, in order to mitigate the damage to the surrounding rock caused by superimposed mining disturbance and to enhance the stability of the surrounding rock in the roadway for safe production, tunneling will be halted when it reaches 40 m in front of the 5201 working face and will resume 80 m behind the advancing working face.

(2)

Zone 2 (80 m after face convergence to the open-off cut of the 4201 working face):

Since this zone is no longer affected by the mining movement of the 5201 working face, the coal-pillar-side anchor cable reinforcement is unnecessary, and the rest of the parameters are consistent with Zone I, as shown in Figure 10b.

(3)

Zone 3 (the 4201 working face’s open-off cut—back 20 m):

The roof and both ribs are supported by bolts with a specification of φ20 mm × 2000 mm, installed at a spacing of 800 mm × 900 mm (transverse × longitudinal).

This zone is influenced by the concentrated load of solid coal on the outer side of the overlying goaf, which means the support needs to be strengthened. The roof is reinforced with cable anchors of φ17.8 mm × 7300 mm, installed at a spacing of 1600 mm × 1800 mm (transverse × longitudinal) in a 3-3-3 arrangement, and both ribs are reinforced with short cable anchors (φ17.8 mm × 4300 mm), installed in pairs per row at a spacing of 1600 mm × 1800 mm (transverse × longitudinal), as shown in Figure 10c.

(4)

Zone 4 (20 m after the 4201 open-off cut—5201 open-off cut):

The roof anchor cable is adjusted to a “2-2-2” arrangement, installed in pairs per row at a spacing of 1600 mm×1800 mm; this cancels the reinforcement of both ribs via the side anchor cable. The rest of the parameters are the same as in Zone 3, as shown in Figure 10d.

(5)

Zone 5 (the 5201 working face open-off cut to the 5202 working face open-off cut):

The roof anchor cable is adjusted to a “2-1-2” arrangement, installed in pairs per row at a spacing of 1600 mm×1800 mm; this cancels the reinforcement of both ribs via the side anchor cable. The rest of the parameters are consistent with those in Zone 3, as shown in Figure 10e.

7. Industrial Tests

In order to test the effect of the zoned support scheme, monitoring stations were deployed in the following five stress influence zones of the 5202 air-return roadway, in order to measure roof-to-floor convergence and rib convergence: Zone 1, influenced by the mining of the adjacent working face under goaf; Zone 2, influenced by the adjacent goaf lateral abutment stress under goaf; Zone 3, influenced by the stress concentration of the overlying solid coal; Zone 4, influenced by the adjacent goaf lateral abutment stress under the overlying solid coal; and Zone 5, influenced by the stabilized stress under the overlying solid coal. The apparatus utilized for measuring the roadway deformation are shown in Figure 11. The monitoring results are shown in Figure 12. The results indicate that, within 20 days of roadway excavation in Zone 1, the surrounding rock’s convergence rate was slower than in other zones. Between days 20 and 30, the convergence stabilized. Between days 30 and 40, the convergence increased significantly, at a higher rate than during the initial 20 days, due to mining-induced influence from the 5201 working face, causing rapid roadway deformation. In the final 20 days, the growth rate decreased notably and approached stability. In Zones 2–5, the surrounding rock’s convergence increased rapidly within 20 days. Roof-to-floor and rib convergence in all zones remained below 400 mm, indicating effective overall deformation control of the surrounding rock. The deformation magnitude ranking is as follows: Zone 3 > Zone 1 > Zone 4 > Zone 2 > Zone 5. The field performance of the 5202 air-return roadway is shown in Figure 13.

8. Conclusions

This study utilized the 5202 air-return roadway in Huaye Coal Mine as its engineering background and as a case study. Numerical simulation, theoretical analysis, and industrial application methods were adopted to analyze the laws of the dynamic evolution of vertical stress in such roadways. In this way, the mine pressure behaviors of AWFDRs in ultra-close coal seams were clarified, thereby enabling the proposal of zoned support technology as a solution, which was subsequently validated through industrial field application.

(1)

The 5202 air-return roadway, acting as the AWFDR in an ultra-close coal seam, exhibits five different characteristic behaviors of mine pressure zones during excavation. The details are as follows: Zone 1 is influenced by the mining of the adjacent working face under goaf; Zone 2 is influenced by the adjacent goaf lateral abutment stress under goaf; Zone 3 is influenced by the stress concentration of the overlying solid coal; Zone 4 is influenced by the adjacent goaf lateral abutment stress under the overlying solid coal; and Zone 5 is influenced by stabilized stress under the overlying solid coal. The mine pressure behaviors of these zones are ranked from most intense to weakest as follows: Zone 3 > Zone 1 > Zone 4 > Zone 2 > Zone 5.

(2)

During the excavation of the AWFDR under goaf, the first stress peak of the roadway’s surrounding rock occurred within the range of 40 m ahead to 80 m behind relative to the advancing longwall face, due to the influence of the mining of the working face (Zone 1). Subsequently, as the roadway was driven into the overlying solid coal, the second stress peak of the roadway’s surrounding rock occurred due to the stress concentration of the overlying solid coal (Zone 3), which was 20% higher than the first stress peak. The theoretical calculation results show that the stress concentration of the overlying solid coal on the lower coal seam’s 5202 air-return roadway was mainly in the range of 0~20 m, and that the influence range was related to the length of the plastic and elastic pressurized areas on the side of the overlying solid coal mining area; additionally, the influence range was about 1.4 times the sum of the length of both of them.

(3)

Taking into account the characteristic behaviors of mine pressure zones as zoning support criteria, a basic support scheme was proposed, which involves using bolt–mesh–beam supports combined with shed supports under the goaf and bolt–mesh–beam supports combined with roof anchor cables under the overlying solid coal. Additionally, in Zones 1 and 3, roof anchor cables or rib anchor cables are supplemented as reinforcing supports, which combine with the basic support scheme described above to form the zoned support scheme for the AWFDR. In this study, roof-to-floor and rib convergence in all zones remained below 400 mm, indicating the effective overall deformation control of the surrounding rock.