Study on the Evolution Characteristics of Surrounding Rock and Differentiated Support Design of Dynamic Pressure Roadway with Double-Roadway Arrangement

Abstract

To elucidate evolutionary characteristics of the surrounding rock failure mechanism in a double-roadway layout, this work is grounded on in the research context of the Jinjitan Coal Mine, focusing on the deformation and failure mechanisms of double roadways. This paper addresses the issue of resource wastage resulting from the excessive dimensions of coal pillars in prior periods by employing a research methodology that integrates theoretical analysis, numerical simulation, and field monitoring to systematically examine the movement characteristics of overlying rock in the working face. On that basis, the size of coal pillar is optimized. The advance’s stress transfer law and deformation distribution characteristics of the return air roadway and transport roadway are studied. The cause of the asymmetric deformation of roadway retention is explained. A differentiated design is conducted on the support parameters of double-roadway bolts and cables under strong dynamic pressure conditions. The study indicates that a 16 m coal pillar results in an 8 m elastic zone at its center, balancing stability with optimal resource extraction. In the basic top-sloping double-block conjugate masonry beam structure, the differing stress levels between the top working face’s transport roadway and the lower working face’s return air roadway are primarily due to the varied placements of key blocks. In the return air roadway, floor heave deformation is managed using locking anchor rods, while roof subsidence is controlled with a constant group of large deformation anchor cables. The displacement of surrounding rock increases under the influence of both leading and lagging pressures from the previous working face, although the change is minimal. There is a significant correlation between roadway deformation and support parameters and coal pillar size. With a 16 m coal pillar, differential support of the double roadway lowers the return air roadway deformation by 30%, which improves the mining rate and effectively controls the deformation of the roadway.

1. Introduction

During the mining process, the double-roadway arrangement can efficiently and rapidly establish the groundwork for the prompt commissioning and utilization of nearby working faces. In recent years, the layout of double roadways has been effectively applied. While bringing considerable economic benefits, it has also produced some engineering problems that cannot be ignored [1,2]. In double-roadway operations, the stress distribution around the transport roadway of the current working face and the return air roadway of the subsequent working face is altered, with each roadway experiencing a different frequency of mining disturbances. The difference in service cycles leads to complex and variable stress in the surrounding rock of the roadway and within the coal pillar between roadways.

The double-roadway heading face layout has been the subject of extensive examination in recent years by numerous academics and researchers. Drawing on extensive monitoring data, Kang et al. [3] studied the stress distribution and deformation/failure characteristics of surrounding rock in multi-roadway configurations, finding that surrounding rock deformation and failure mainly occurred in the goaf area behind the working face. Yi et al. [4] proposed a unified nonlinear strength theory for rocks and studied the evolution and failure of “butterfly-shaped” plastic zones in deep tunnels to reference underground resource development, while Lei et al. [5] thereafter proposed differentiated support measures combining plastic zone scope and development patterns to effectively control tunnel deformation. Zhu et al. [6] investigated the factors contributing to wall panel deformation and floor heave in deep mining roadways, proposing a control strategy focused on one roof and two corners, which successfully mitigated roadway deformation. As the depth of the working face gradually increases, the mining process causes disturbance in the original rock and soil body [7], which reduces the effect of traditional support technology gradually [8]. Jin et al. [9] proposed an asymmetric technology for a dynamic pressure roadway based on the Guotun coal mine. Wu et al. [10] used a POS-SVM prediction model to reasonably select the coal pillar size; Tian et al. [11] analyzed the support failure of narrow coal pillar roadways in double-roadway excavation and proposed a drilling pressure-relief technology to release stress for resolving deformation instability, while Liu et al. [12] simulated coal pillar failure patterns in double-roadway layouts via UDEC and applied composite support to reduce fracture zones and enhance stability. Serhii et al. [13] developed the Smart 4 controller and its supporting installation module and verified the applicability and effectiveness of monitoring the vibration load of drilling tools through industrial field tests. Peng et al. [14,15] investigated control techniques for roadway-surrounding rock in the longitudinal mining of thick coal seams, employing partition support to address varying stress levels. Wu et al. [16] uncovered the mechanical mechanism of surrounding rock failure in gently inclined, closely spaced coal seam extraction roadways, proposing a stage-specific collaborative control scheme. Guo [17] proposed an NPR anchor cable, multi-dimensional grouting, and roof-cutting pressure relief for deep soft-rock roadway deformation control based on stress compensation theory, while Cheng et al. [18] presented roof-cutting pressure relief plus grouting reinforcement for water-bearing soft-rock roadway swelling floor heave. Peng et al. [19,20] investigated a reinforcement strategy for a 12 m wide coal pillar, suggesting that NPR constant-resistance large-deformation anchor cables and pressure anchor cables could effectively reduce its deformation.

Numerous scholars, including Guo et al., have extensively researched the stability of double-entry layout coal pillars and the failure mechanisms of surrounding rock in roadways [21] and investigated the movement and fracture patterns of overlying strata above thick coal seams with hard roofs. Liu et al. [22] obtained the variation laws of the principal stress and dilatancy characteristics in the plastic zone of retained entries, explaining the causes of the asymmetric deformation. Kumar et al. [23] calibrated field studies with the CMRI empirical formula to evaluate pillar performance under deeper overburdens. Sun et al. [24] revealed the roof evolution laws under repeated mining: primary mining formed an “F-shaped” overlying rock structure, while secondary mining produced a “T-shaped asymmetric/symmetric structure”. Yang et al. [25] proposed a new pillar spalling model accounting for gradually reduced pillar sizes and the confining behavior of spalled coal fragments. Wang et al. [26] analyzed the stress and displacement evolution mechanisms under varying coal pillar widths. Tao et al. [27] identified mining-induced stress as the primary cause of roadway floor deformation and instability, proposing methods to improve stress conditions via optimized protective coal pillar widths for surrounding rock control. Chen et al. [28] studied the double-entry layout beneath upper-seam pillars with interlayer spacing below 2 m. Yang et al. [29] developed a differentiated control technology for double entries based on rib anchor cables and floor corner bolts. Hao et al. [30] investigated the instability mechanisms and key influencing factors of deformation failures such as roof falls, rib convergence, and floor heave in return air roadways. Xue et al. [31] investigated the failure characteristics of surrounding rock during goaf entry excavation. Gu et al. [32] identified large asymmetric deformation unevenly distributed along the roadway axial direction as a characteristic of surrounding rock deformation, proposing an asymmetric control strategy including rock homogenization, mass reinforcement, and coupled support in key zones. Li et al. [33] attributed intense non-uniform roadway subsidence and anchor cable fracture to expansion pressure and fragmentation/swelling deformation corresponding to rock failure within heterogeneous plastic zones.

The above scholars have carried out extensive research on roadway support and surrounding rock control. Double-lane synchronous tunneling has great advantages in coal mine production, which can significantly improve resource recovery rate and mining efficiency. Integrating the actual conditions of the double-lane layout at the Jinjitan Coal Mine with previous research findings, in this paper, theoretical calculations, numerical simulations, and on-site monitoring are employed to optimize the surplus coal pillar design of the double-roadway layout working face. Based on the different stress characteristics during the excavation of the 16 m coal pillar double roadway, a differential support design is proposed for the two types of roadways. It solves the superimposed damage due to the rotation of the two-sided working face in the return air roadway. At the same time, it can ensure the stability of the surrounding rock. The design’s rationality is confirmed through on-site monitoring, providing a reliable safeguard for the safe and efficient operation of the working face.

2. Project Overview

Geological Conditions

In the Yushen Mining Area of Shaanxi Province, China, lies the Jinjitan Coal Mine; the minefield has a strike length of about 11.44 km, an inclination width of about 8.77 km, an area of about 91.6206 km², and an approved production capacity of 17 million t/a. The 113 working face, a fully mechanized caving face, is situated in the eastern wing of the mine’s initial panel. It features a gradient of 300 m, a strike length of 4847.6 m, and a coal seam thickness between 7.9 and 12.6 m, averaging 9.50 m. The coal seam dips at an average angle of less than 1° and has a straightforward geological structure. The drilling data provide evidence that no significant faults are present at the cut hole. Figure 1 depicts the drilling columnar structure, the mining plane layout of the working face, and the geographical location map of the Jinjitan Coal Mine [34].

3. Size Optimization of Coal Pillar in Double-Roadway Excavation

The establishment of an appropriate coal pillar width between the two roadways is essential for ensuring the safe extraction of resources in a big-mining-height double-roadway configuration’s working face. Effectively preventing associated incidents can be achieved by integrating the optimal coal pillar width, optimizing roadway support parameters and construction techniques, and monitoring surrounding rock deformation [35,36,37].

3.1. Theoretical Calculation of Coal Pillar Width

The dimension L of the coal pillar separating roadways is expressed as $\mathrm{L}={\mathrm{L}}_{\mathrm{A}}+{\mathrm{L}}_{\mathrm{B}}+{\mathrm{L}}_{\mathrm{C}}$. In the formula, ${\mathrm{L}}_{\mathrm{A}}$ represents the length of the crushed zone in the inter-entry coal pillar; ${\mathrm{L}}_{\mathrm{B}}$ denotes the length of the relatively stable zone in the inter-panel coal pillar; and ${\mathrm{L}}_{\mathrm{C}}$ signifies the length of the bolt support zone, as shown in Figure 2.

Once the primary mining influence stabilizes, the boundary of the coal pillar between areas A and B reaches a limit equilibrium state, satisfying the following equation:

$$\left\{\begin{array}{c}{\displaystyle \frac{{\partial \sigma }_x}{{\partial }_x}}+{\displaystyle \frac{{\partial }_{\tau }}{{\partial }_y}}+f_x=0\\ {\displaystyle \frac{{\partial \sigma }_y}{{\partial }_y}}+{\displaystyle \frac{\partial {\tau }_{xy}}{{\partial }_x}}+f_y=0\\ { \tau }_{xy}=-\left({\sigma }_y\tan \phi +C\right)\end{array}\right.$$

(1)

In the equation, ${\sigma }_x$ and ${\sigma }_y$ represent the principal stresses in the x and y directions. $\tau$ is the shear stress; $f_x$ and $f_y$ are the friction coefficients in the x and y directions; $\phi$ is the coal seam interface’s internal friction angle; $C$ is the cohesion of the coal seam interface.

Through analysis and calculation, the width of Area A within the coal pillar is determined to be:

$$L_A={\displaystyle \frac{m\lambda }{2\mathrm{t}\mathrm{a}\mathrm{n} \phi }}\ln \left({\displaystyle \frac{K\rho gH+\frac{C}{\tan \phi }}{\frac{C}{\tan \phi }+\frac{P_x}{\lambda }}}\right)$$

(2)

In the formula, the extraction elevation of the coal seam is denoted by $m=5.8$; $\lambda$, the lateral pressure coefficient, from on-site monitoring, is $\lambda =0.316$; $K$, the stress concentration factor, is $K=2$; $\rho$, the average density of the overlying strata, is $\rho =2.5 \mathrm{t}/{\mathrm{m}}^3$; $\mathrm{H}$, the burial depth of the roadway, is 263 m; $C$, the cohesion of the coal seam interface, is $C=2.82 \mathrm{M}\mathrm{P}\mathrm{a}$; $\phi$, the internal friction angle of the coal seam interface, is $\phi =37.64°$; and $P_x$, the lateral pressure of the support on the roadway, is 0.15 MPa.

Substituting each parameter into the equation yields ${\mathrm{L}}_{\mathrm{A}}$ = 7.12 m.

The bolt support length of the coal roadway is 2.0 m, and the width of the relatively stable central region can be calculated using Equation (3):

$${\mathrm{L}}_{\mathrm{B}}=\xi \left({\mathrm{L}}_{\mathrm{A}}+{\mathrm{L}}_{\mathrm{C}}\right)$$

(3)

In the formula, $\xi$ is the stability coefficient, set to 0.7, and L_B = 6.384 m.

Substituting each parameter, we get L = 15.504 m, that is, through a theoretical analysis, the suitable width for the section of the coal pillar is 16.00 m.

3.2. Coal Pillar Optimization Design Simulation Calculation

3.2.1. Model Establishment

Based on the relevant geological conditions of the Jin Jitan Coal Mine and the borehole histogram of rock strata, the FLAC3D 6.0 numerical model software was used to establish the basic model of the 113 and 111 mining faces and roadway coal pillar, and the associated physical characteristics of various rock layers were allocated [38].

Table 1. Mechanical parameters of rocks.

Lithology	Density/(kg/m3)	Elastic Modulus/(GPa)	Tensile Strength/(MPa)	Cohesion/(MPa)	Internal Friction Angle/(°)
Medium sandstone	2653	12.5	4.6	2.5	31
Mudstone	2589	8.67	2.55	1.7	25
Coarse sandstone	2813	12.35	6.5	2.7	20
Siltstone	2680	16.5	6.97	1.3	21
Coal	1390	1.68	1.1	1.2	19
Thickness	2531	14.6	6.58	2.0	32

details the physical parameters of the respective rock layers. The model used the Mohr–Coulomb constitutive model. The engineering geological survey report shows that the dip angle of the coal seam and overlying strata in the working face is less than 1°. The mechanical effect, parameter change, and calculation error caused by it are all in the range that can be ignored in engineering, and the influence on the simulation results is usually negligible. This is not significantly different from the treatment method that regards the dip angle as 0°. Therefore, in order to facilitate the simulation calculation and improve the calculation efficiency, the model was set to 0°, with constrained boundaries. The model simulated a free plane at the top, applying a downward stress of 4.8 MPa and a gravitational acceleration of 9.8 m/s² to represent the overlying strata. The model measured 680 × 500 × 150 m, as illustrated in Figure 3.

3.2.2. Computational Simulation

The study modeled the plastic zone state and stress distribution in the surrounding rock for varying coal pillar widths of 12, 14, 16, and 18 m after coal seam mining. During the simulation, unloading the mining coal seam at the working face released the surrounding rock’s stress field, leading to significant variations in the stress field adjacent to the coal pillar. The medium-high stress zones from mining faces 119 and 111 were interconnected, resulting in stress superposition in the high-stress area. When the coal pillar was 12 m tall, it occupied the medium-high stress core area, with stress exceeding 40 MPa. Therefore, a coal pillar width of less than 12 m negatively impacted its stability. This may lead to the edge of the coal pillar peeling off and the interior being crushed, which in turn may cause the overlying rock layer to collapse, seriously affecting the safe production of the working face. As the dimensions of the coal pillar expanded from 12 m to 14 m, the stress peak area transitioned from a singular zone to two clearly distinguishable zones. As the stress peak decreased and the coal pillar’s stability improved, a 2–8 m stress slow-drop zone developed at the pillar’s center. When the coal pillar size ranged from 14 to 16 m, the mutual influence of the two stress peak areas gradually became weak, and a low-stress area appeared in the middle, with a large elastic range, as shown in Figure 4.

The distribution of the plastic zone indicated that as the coal pillar size exceeded 16 m, both the stability and the elastic zone range increased. In comparison, the 16 m coal pillar not only ensured the long-term stability of the surrounding rock by forming a stable low-stress elastic zone but also kept coal resource losses within a reasonable range, achieving the optimal balance between safe mining and efficient resource utilization. As shown in Figure 5, coal resources would be wasted. As a result, keeping the coal pillar’s size around 16 m is preferable.

A station was installed in that section to monitor the lateral abutment pressure of the coal pillar. The collected data were sorted and fitted to derive the stress distribution curve for coal pillars of varying sizes, as illustrated in Figure 6.

The stress curve indicated that a 12 m section coal pillar reached its vertical stress peak approximately 5 m inward from the edge. With a slight increase in distance, the vertical stress in the coal pillar remained nearly constant. This suggests that in a 12 m section coal pillar, most plastic deformation occurs within the pillar, while the elastic core area is minimal. As the section of the coal pillar expanded to 14 m, there was an obvious elastic core area; the sustaining pressure inside the coal pillar dropped significantly when the size of the section coal pillar approached 16 m, giving rise to an elastic core area of approximately 6 m. An 18 m coal pillar size enhancec the stability of the section by expanding the elastic core area, though it also led to increased resource wastage. This did not facilitate the enhancement of coal resource recovery rates.

4. Double-Roadway Layout’s Surrounding Rock Movement Law

When the two roadways are arranged, the excavation time of the two roadways is the same, but the service life is different, and the degree of mining influence on both sides is different. New maintenance challenges will emerge for the two routes due to the transport roadway of the previous working face being affected by the double-roadway excavation and the advancing support pressure of the current working face. Similarly, the return air roadway of the subsequent working face is influenced by tunneling, the advancing effect of the working face, and the dynamic impact of mining at the preceding working face.

4.1. Characteristics of Rock Mass Movement

During the initial excavation of the double roadway, the vertical stress in the return air roadway for the subsequent working face is similar to the transport roadway’s stress from the previous working face. The maximum vertical stress reaches 10.4 MPa, with the stress distribution exhibiting a left–right symmetrical pattern along the coal pillar’s central line. The convergence values on both sides are essentially identical, with the roadway’s maximum horizontal deformation approximating 59 mm, as depicted in Figure 7.

During back extraction on one side of the double-roadway working face, the collapse of the overlying rock in the goaf leads to the formation of “O-X” fractures due to the periodic pressure of the main roof. This results in a masonry beam structure formed by rock blocks following the main roof’s periodic fractures along the strike and dip directions of the working face [39], as shown in Figure 8.

The overlying strata of the double roadway interact with the roadways via the immediate roof above the roadways. The caving behavior of overlying strata after coal seam extraction at the working face is influenced by the fracture patterns and collapse state of primary roof layers. Consequently, a comprehensive surrounding rock structure for double roadways is suggested, encompassing top coal, direct roof, main roof, and the load rock stratum impacting the main roof. This model is called double-roadway conjugate rock masonry beam model. A2 and B2 are triangular rock blocks located on the goaf side of the working face. In a working face strike, rock block A3 limits the condition of rock block A2, which has collapsed into the goaf, while rock block B2 restricts the movement of rock block B3. In the dip direction, rock block B3 constrains the movement of rock block B2. The state of the main roof affects the stability of the double roadway, and the state after the main roof is broken directly forms the top boundary of the double roadway. The fundamental reason for the difference in forces on the transport roadway of the upper working face and the return air roadway of the subsequent working face is the varying position of key block B2.

Structurally, the ends of the working face break into triangular blocks, exerting vertical and lateral pressure on the roadway due to the combined effects of the overlying strata and these blocks. Furthermore, in terms of temporal factors, the service duration of the preceding working face’s transport roadway spans only one mining cycle of the current working face, whereas service duration of the subsequent working face’s return air roadway covers two mining cycles of the current working face and the next two working faces—doubling that of the preceding transport roadway. The transport roadway of the upper working face is mainly affected by tunneling and the advance’s abutment pressure of the working face. In contrast, the return air roadway of the subsequent working face is influenced not only by tunneling and the current working face’s advance’s impact but also by periodic disturbances from mining activities in the previous working face’s goaf.

Therefore, the primary cause of the loading difference between two roadways is the different location of key block B2 in the transport roadway of the upper working face compared to the return air roadway of the succeeding working face.

4.2. Characteristics of Differential Stress in Surrounding Rock

The overall vertical stress of the coal pillar is forced to change from a merged state to a diverted state during the double roadway’s excavation and mining. A stress peak was observed at the coal pillar 50 m behind the working face, displaying an asymmetric pattern characterized by two peaks and one valley. The coal pillar next to the return air roadway had a significantly higher vertical stress than that of the transport roadway. When the working face was 40 m ahead, the coal pillar’s vertical stress tended to stabilize as it declined, exhibiting a symmetrical pattern. Vertical stresses in both the shallow and deep sections of the return air roadway and transport roadway roofs progressively rose. In the 0–7 m range, the return air roadway’s roof endured greater vertical stress than the transport roadway, whereas in the 7–16 m range, the transport roadway’s ceiling experienced slightly higher vertical stress compared to that of the return air roadway, as seen in Figure 9.

4.3. Surface Proximity Differential Characteristics

During mining, on-site monitoring is performed on the roof, floor, and sides of the double roadway, as illustrated in Figure 10. The results show that due to the influence of mining disturbance, the roof subsidence of return air roadway and transport roadway decreases with the increase in the distance of the leading working face, and the attenuation rate gradually decreases. At the working face position, the convergence of the return air roadway is about 50% of that of the transport roadway. The subsidence of the return air roadway is consistently less than that of the transport roadway, and this difference diminishes as the working face distance increases.

The floor heave in both the return air roadways and transport roadways diminishes with increased distance from the primary working face. The transport roadway’s bottom plate shows significant fluctuation, with a maximum bulge of 415 mm at the working face, decreasing sharply to a minimum of 0.23 m at the 50 m advance, as shown in Figure 11. As the distance from the working face increases, the floor heave’s size gradually decreases. The floor heave of the return air roadway reaches a peak of 324 mm at 60 m behind the goaf then gradually decreases and stabilizes 40 m ahead of the working face. The bottom drum volume of the return air roadway exceeds that of the transport roadway, with the volume difference diminishing as the distance to the working face increases.

5. Double-Roadway Differential Support Design and Application

5.1. Principle of Differential Control

A differential support technology for the roadway-surrounding rock is proposed, considering the varying stress characteristics encountered during the excavation of a 16 m coal pillar double roadway. The roadway was divided into two categories, one was the 109 transport roadway (class A roadway), which only experienced one mining, the other was the 111 return air roadway (class B roadway), which needed to undergo repeated mining.

In dynamic pressure roads, to manage the widespread deformation of the surrounding rock, the differential support method is applied [40], so as to ensure that the Class A roadway does not need to be repaired over a large area during use; the deformation controllability of Class B roadways under the influence of mining-induced dynamic pressure from adjacent working faces ensures their capacity to support the rapid advancement of subsequent working faces, maintains the surrounding rock’s stability in the service cycle of roadway, and decreases the rate of road renovation and production costs while achieving the safe and effective return of coal resources.

5.2. Differentiated Control Scheme

5.2.1. Differentiated Support Scheme for Class A Roadways

Threaded steel bolts measuring Φ 20 × 2000 mm were used for the roof of the return air roadway and the non-minable side, whereas glass fiber-reinforced plastic bolts of the same dimensions were employed for the minable side. Roof bolts were Φ 20 × 2000 mm threaded steel bolts with a spacing of 1000 mm × 1000 mm. A Φ 6.5–100 × 100 mm steel mesh was laid on the roof and non-minable side. The roof was reinforced with anchor cables, using Φ 17.8 × 8000 mm anchor cables positioned along the roadway centerline, and each row contained two cables with a row spacing of 200 × 5000 mm. For roadway side reinforcement, anchor ropes were installed vertically on the coal side at 800 mm and 2300 mm from the roof. These Φ 17.8 × 4500 mm anchor ropes had a row spacing of 5000 mm, with the bottom row offset 2500 mm along the strike direction from the top row, forming a staggered triangular pattern. The floor corner bolts at the working face measured Φ 20 × 2000 mm (threaded steel) at a 45° angle, while the pillar-side floor corner bolts also measured Φ 20 × 2000 mm (threaded steel) at a 45° angle. The support design is illustrated in Figure 12.

5.2.2. Differentiated Support Scheme for Class B Roadways

The roof of the return air roadway and non-minable side utilized Φ 20 × 2000 mm threaded steel bolts, whereas the minable side employed Φ 20 × 2000 mm glass-fiber-reinforced plastic bolts. The roof bolts were spaced 1000 mm apart in both directions, while the side bolts had a spacing of 1200 mm by 1000 mm. A Φ 4.5–100×100 mm steel mesh was installed on the roof and non-minable side. Anchor ropes were employed for roof reinforcement, with Φ 17.8 × 8000 mm anchor ropes selected. The anchor ropes were arranged along the roadway centerline, one per row at 5000 mm intervals. The coal pillar’s side bottom corner bolt measured Φ 20 × 2000 mm (threaded steel) at a 45° angle, as shown in Figure 13.

5.2.3. Actual Usage Effect of Class B Roadway Support

Class B roadway (111 return airway) achieved stable control of the surrounding rock through differentiated support design after being subjected to the dual mining impacts of the previous working face (109) and the current working face. The floor showed no significant bulging or cracks, the coal bodies on both sides were intact, the roadway profile was regular, and no asymmetric deformation occurred, confirming the adaptability of differentiated support to complex stress environments. The roof surface showed no significant subsidence or cracks, and the coal bodies in the areas covered by steel mesh were intact, demonstrating the dynamic control effect of constant-resistance anchor cables on roof subsidence, as shown in Figure 14.

6. On-Site Data Monitoring

6.1. Measurement Point Arrangement

The mining of the 109 working face significantly affected the return air roadway of the 111 working face, as indicated by the mine pressure characteristics in previous mining roadways. To effectively assess the support efficacy of the differentiated roadway during the extraction of the 109 working face, the return air roadway of the 111 working face was monitored. The station layout points are illustrated in Figure 15.

Three anchor stress stations were installed in the 111 return air roadway at intervals of 100 m, 200 m, and 300 m from the working face. The load force of the anchors was specifically monitored at a distance of 100 m from the working face. The cross-measuring point method monitored the convergence of the roadway roof and sides within 20 m before and after the working face. A 10 m deep roof borehole was drilled 15 m ahead of the working face between No. 1 and No. 2 stations for inspection purposes and detection of roadway support construction quality.

6.2. Analysis of Data Monitoring Results

6.2.1. Monitoring of Anchor (Bolt) Forces in the Return Air Roadway

Each station was equipped with six stress sensors: one anchor cable was installed on the roof, and one anchor bolt was positioned there as well, one anchor cable was arranged on both sides, and one anchor bolt was arranged on both sides. Figure 16 displays the monitoring results.

Monitoring results indicated that as the working face progressed, the stress on the bolt and anchor cable consistently increased. There was a significant rise in stress 40 m from working face. The peaks at the working face’s reaching station were 90 kN and 300 kN, respectively. The monitoring station moved into the goaf area as the working face passed it, and strain on the anchor bolts (cables) progressively lessened. The stress on anchor cables in the deeper rock mass remained elevated due to the influence of overlying rock layers in the goaf area, compared to when the working face was ahead. Within the 200 m monitored range, the anchor stress initially rose and subsequently declined. The advance support pressure affected an area extending approximately 40 m from the working face, while the rear goaf influenced an area approximately 50 m behind it. The anchor (bolt) force remained within a reasonable range.

6.2.2. Monitoring of Surrounding Rock Convergence in Return Air Roadway

As indicated by the monitoring data, refer to Figure 15 for details, the differentiated support method resulted in minimal roof and floor convergence when the working face was 60 m ahead. Then, as the working face advanced toward the roadway, it continued to rise until the working face was 40 m past the roadway, at which point the roof-to-floor convergence rate remained relatively high. The convergence of the surrounding rock initially increased and then stabilized at a distance of 50 m behind the working face. The maximum convergence between the roof and floor was approximately 486 mm, with the roof subsiding by 192 mm, and the floor heave was about 294 mm [41], which was reduced by about 30%. Typically, the deformation of the roadway’s surrounding rock remained under 10% of its size, maintaining good integrity, as illustrated in Figure 17.

6.2.3. Advanced Working Face Roof Peeping

As the 109 working face advanced 15 m away from the roof peephole, a borehole peeper was employed to observe fracture development in the 10 m borehole located in the roof of the 111 working face’s return air roadway. Figure 18 presents the borehole peep results.

In the range of 0–1 m from the orifice, the coal body on the roof was broken, coal body fractures were more extensively developed up to 2 m from the orifice, the coal body was complete in the range of 3–8 m from the orifice and deeper, and no obvious fracture was found in the borehole. This shows that the roof of the return air roadway can basically remain intact after being disturbed by mining.

7. Conclusions

(1) The layout of double-roadway excavation can guarantee the rapid production layout of the next working face; therefore, the design of roads and coal pillars is especially crucial. In this paper, the dimensions of the pressure-bearing coal pillar were optimized using theoretical calculations and numerical simulations, affirming the practicality of constructing a coal pillar 16 m wide, which achieved the optimal balance between safety and economy.

(2) Numerical simulations were employed to analyze the differential characteristics of surrounding rock damage during tunneling and mining. In the early stage of tunneling, roadways and coal pillars had a balanced and minimal stress distribution. In the mining of the 109 working face, coal pillar stress exhibited an asymmetric “saddle” shape. Notably, the vertical stress in the return air roadway behind the working face was considerably higher than that in the transport roadway of the preceding working face, with significant fluctuations observed at the floor’s two corners.

(3) Considering the distinct stress and strain characteristics of the two roadways, the proposed assistance plan for the return air roadway involved using a continuous resistance, high-deformation anchor cable with a negative Poisson’s ratio at the apex to manage roof rock movement, while anchor rods were applied at the floor’s edges to mitigate floor heave.

(4) Through on-site industrial monitoring experiments, the design of a 16 m coal pillar and differentiated support ensured that the force on the anchor bolts (cables) remained within a reasonable range, reduced the displacement of the roof and floor by approximately 30%, effectively controlled the deformation of the roadway, and ensured the stability of the return airway during the third disturbance.

(5) In the model construction, the MohrcCoulomb elastoplastic model had certain advantages, but there were limitations. In the process of cumulative damage failure, due to the lack of damage cumulative evolution equation, the criterion could not quantify the dynamic process from initial damage to complete damage. In the future, research on the dynamic support design method of strong dynamic pressure roadway will be carried out.