Transferring Pressure Mechanism Across Gob-Side Roadway Goaf with Coal Pillar During Distant Face Mining: A Case Study

Abstract

The gob-side roadway technique is extensively utilized in coal extraction due to its capacity to enhance coal resource recovery efficiency and mitigate mining sequence conflicts. Nevertheless, increasing mining depths lead to progressively intricate stress conditions, posing challenges for maintaining gob-adjacent roadway surrounding rock stability. Taking the belt haulage roadway 1513 (BHR 1513) at Xinyi Coal Mine as an engineering case, this research investigates the application of narrow-pillar gob-side roadway construction under remote working face mining conditions. By integrating field observations, analytical modeling, and computational simulations, the cross-goaf pressure transfer phenomenon and its formation mechanism in narrow-pillar roadways under distant mining operations are systematically examined. Key findings reveal that during the alternating extraction of wide and narrow working faces, the caving angle terminates roof collapse within the narrow working face goaf at the second key stratum (KS2). The subsequent mining of the adjacent wide working face induces stress accumulation in the overlying “T”-shaped strata zone, triggering the instability of the inter-working face island pillar. This pillar failure merges the two goafs into an expanded void, initiating sequential fracture, collapse, and rotational displacement across all overlying key strata (KS). Consequently, previously intact KS above the narrow working face goaf undergo fracturing and rotation, amplifying lateral main roof block subsidence toward the goaf. This kinematic process generates substantial deformation in the narrow-pillar gob-side roadway.

1. Introduction

As a specialized configuration for longwall mining roadways, gob-side excavation technology significantly enhances coal recovery efficiency while resolving mining sequence coordination challenges, thereby constituting a pivotal technique for sustainable coal extraction [1,2,3]. Nevertheless, progressive increases in operational depth and accelerated mining cycles intensify stress field complexity during gob-side roadway development [4,5,6]. This geological evolution necessitates excavating lower sublevel working face roadways prior to complete extraction of adjacent panels. Crucially, dynamic stress interactions from neighboring mining operations substantially compromise the integrity of gob-proximate roadway surrounding rock masses.

Extensive investigations by Chinese researchers have focused on mining-induced disturbances from adjacent working faces to gob-side roadways. Zhang et al. systematically analyzed dynamic pressure evolution during fracturing, rotational displacement, and stabilization of lateral roof structures in small-coal-pillar gob-side roadways adjacent to advancing faces. Their work established a pretension-coupled reinforcement methodology integrating bolt-cable synergy [7]. Concurrently, Yu et al. quantified time-dependent strata movement effects on roadway deformation under progressive mining conditions, developing adaptive cross-sectional control strategies through mechanical–temporal coupling models [8].

Bai et al. investigated stress distribution and failure mechanisms in roof structures of adjacent gob-side roadways, developing a dynamic mechanical model that identifies the roof center as the critical support area [9]. Wang et al. developed a methodological framework for optimizing coal pillar dimensions in gob-side roadways adjacent to advancing faces, establishing integrated control strategies for pillar-roof synergistic stabilization [10]. Liu et al. characterized the full-cycle stress redistribution patterns within thick-seam deep mining environments, deriving critical thresholds for pillar width determination through elastoplastic mechanical modeling [11]. Ma et al. formulated spatiotemporal coordination criteria for parallel roadway development and adjacent mining operations, specifying the following: ① minimum 136 m separation distance with ≥36-day interval for codirectional excavation; and ② <50 m overlap distance when opposing excavation-mining directions cross the stopping line [12]. Regarding surrounding rock control, Gou simulated roadway stability near advancing faces and designed a support scheme that effectively mitigates deformation [13]. Zhao et al. emphasized spatiotemporal coordination as key to managing roadways adjacent to active mining fronts, advocating reinforced roof-pillar systems and initial stabilization measures [14].

Extensive investigations have focused on stress disturbance mechanisms and mitigation strategies for narrow-pillar gob-side roadways under proximal working face mining, establishing critical theoretical frameworks for managing complex surrounding rock conditions. In contrast, the stress transfer patterns and dynamic responses of small-coal-pillar roadways subjected to remote panel mining remain understudied, posing significant knowledge gaps in deep mining geomechanics. This study aims to clarify the mechanism of transferring pressure across goaf in distant mining conditions and propose stability control strategies for gob-side roadways with narrow pillars. Aligned with China’s ‘Guidelines for Safe and Efficient Coal Mining’ and the national ‘Dual Carbon’ goals, this research addresses critical challenges in reducing coal pillar waste and minimizing ground subsidence. By optimizing pillar design and support systems, it contributes to sustainable resource utilization and environmentally responsible mining practices. With the engineering background of the BHR 1513 of the Xinyi Coal Mine, this paper introduces the engineering case of gob-side roadway driving with a small coal pillar in the distant face mining and discovers the transferring pressure across the goaf effect and occurrence mechanism of a small-coal-pillar roadway under the condition of such distant face mining. Through theoretical analysis, the occurrence mechanism of transferring pressure across the goaf effect of the gob-side roadway is clarified, that is, the island coal pillar between two goafs collapses as the result of strata stress in the upper “T” type area, connecting the two goafs to form a larger goaf, which severely deforms the gob-side roadway. The results of this research can provide a reference for mining design.

2. Engineering Overview

2.1. Basic Conditions of Roadway

This study investigates the 3up coal seam at the Xinyi Coal Mine (Jining, Shandong, China), with an average working face burial depth of 507 m. To minimize surface subsidence impacts on adjacent villages, the mine adopted alternating wide-narrow working face mining during initial operations (Figure 1). Following the extraction of the 60 m wide strip working face 1513, a 5 m wide coal pillar was preserved along the goaf boundary to facilitate the construction of BHR 1513—a 1008 m fully gob-side roadway serving working face 1513. Concurrently, adjacent working face 1511 (200 m width) underwent active extraction, classifying BHR 1513 as a narrow-pillar gob-side roadway under distant mining conditions.

The main coal seam of the working face 1513 is a 3up coal seam, with a Platts hardness f = 1~2. The average thickness of the coal seam is 3.11 m, with siltstone (1.6 m) as the direct roof and fine sandstone (3.6 m) as the main roof. The other strata of the roof are mainly mudstone and sandstone, and the floor is mainly siltstone and fine sandstone. Based on the distinguishing of the occurrence of KS in the strata in the literature [15], the KS of the roof are further studied and distinguished to KS7, and the KS in the overlying strata of the roof of the coal seam of the working face 1513 are determined to be KS1, KS2, KS3, KS4, KS5, KS6, and KS7 from bottom to top.

2.2. Roadway Maintenance Effect

Prior to BHR 1513 test roadway excavation, a 5 m residual coal pillar was preserved along the goaf periphery of strip working face 1509 within TTR 1511 for gob-side construction (Figure 1). During TTR 1511 development, the active extraction of working face 1507 (60 m offset from goaf boundary) created analogous geomechanical conditions between TTR 1511 and BHR 1513. This geological consistency renders the systematic analysis of TTR 1511’s support design and deformation mechanisms critical for optimizing BHR 1513 surrounding rock control strategies.

The initiation of the TTR 1511 driving commenced two years subsequent to the conclusion of the strip working face 1509 mining. This driving took place adjacent to the progressing working face 1507. Prior to the commencement of working face 1507 mining, the roof movements within the goaf of the strip working face 1509 had already stabilized, accompanied by a stationary mining-induced stress field surrounding the goaf. As a result, the TTR 1511 was positioned in a comparatively stable stress environment, unaffected by fluctuating stresses. The encounter between the track roadway working face 1511 and the mining working face 1507, spanning from a convergence of 50 m to a staggering distance of 200 m, would induce stress disturbances subsequent to the mining of working face 1507. These disturbances manifested as significant roof inclination and deformation, along with overall displacement of the coal pillar sidewalls within the roadway (Figure 2).

2.3. Analysis of Cause of Deformation and Instability of Surrounding Rock of Roadway

Before mining the working face 1507, the TTR 1511 is in a stable lateral residual abutment stress zone and free from stress fluctuation. After the working face 1507 becomes a goaf upon completion of mining, there exist island coal pillars between the strip working face 1509 and the working face 1507, which collapse and become unstable due to increased stress after their formation in the goaf of the working face 1507, causing the breaking of high overlying strata of the strip goaf 1509, affecting the below TTR 1511 and causing severe deformation of the roadway. The mine pressure appearance that the mining of the working face poses dynamic pressure influence on the gob-side roadway one goaf away is called “transfer pressure across goaf”.

3. Occurrence Mechanism of Transferring Pressure Across Goaf of Alternative Wide and Narrow Working Faces

3.1. Principle of Transferring Pressure Across Goaf Effect of Alternative Mining of Wide and Narrow Working Faces

Taking the BHR 1513 as an example, this section introduces the occurrence mechanism of transferring pressure across the goaf of alternative wide and narrow working faces. After mining the coal seam, the overlying strata in the goaf breaks upwards layer by layer starting from the caving angle, and the final caving height of the overlying strata roof is related to the width of the working face. According to the method in Han et al., used to distinguish the span and limit span of the KS of the overlying strata based on the width of the working face [16], it is determined that when the strip working face 1513 with a width of only 60 m becomes a goaf upon the completion of mining, only KS1 in the KS of the overlying strata breaks, with no breaking of other KS yet. When the adjacent working face 1511 becomes a goaf upon the completion of mining, the two goafs connect to form a goaf with a width of 265 m, breaking all 7 KS in the overlying strata.

The study by Yang et al. shows that prior to the mining of the working face 1511 [15], coal pillar A only needs to bear the load of the “T”-shaped structure below KS2 as only KS1 in the overlying strata breaks and collapses. As the working face 1511 becomes a goaf with a width of 200 m with the gradual mining, there is an increase in the “T”-shaped strata that the coal pillar A needs to bear, causing increasing stress for the coal pillar. Coal pillar B finally collapses with the increase in the stress, thus the goaf 1511 and the strip goaf 1513 connect to form a larger goaf. The roof overlying strata continues to break and collapse upwards, and all 7 KS break. As a result, the original unbroken KS above the strip goaf 1513 break and rotate to form dynamic pressure rock blocks, causing stress disturbance to BHR 1513 and a larger roadway deformation.

3.2. Analysis of Disturbance Mechanism of Transferring Pressure Across Goaf Effect to Gob-Side Roadway

The lateral roof fracture in the goaf upon the completion of the mining of the strip working face 1513 is shown in Figure 3a; meanwhile, the position of gob-side roadway driving with the small coal pillar is usually located in a low-value stress area below block A and block B [17,18,19]. During the driving of the BHR 1513, the disturbance to its overlying strata will not affect the stability of block A and B. At this time, no changes are observed in the deformation and mechanical characteristic of block A and B, which bear the weight of the collapsed strata of the overlying strata to shield the roadway. During driving, the deformation of the surrounding rock is mainly caused by the stress concentration of the broken coal rock mass within the lateral low-value stress zone in the goaf due to the driving disturbance.

Before mining the working face 1511, only KS1 in the roof overlying strata of the strip goaf 1513 breaks and collapses, while high KS do not break. Therefore, block A and B only need to bear the weight of the strata below the unbroken KS2. The rotation deflection of block A and B at this stage are relatively small as it is related to the weight of the strata that it bears (Figure 3a). After the completion of the mining of the working face 1511, the two goafs connect, causing the high KS in the roof of the overlying strata of the strip goaf 1513 to break. At this time, block A and B further rotate and subside towards the strip goaf 1513 due to increased load-bearing. Due to the rotation and subsidence of block A and B, the roof of the roadway further tilts towards the goaf, and the coal pillar sidewall squeezes into the roadway space (Figure 3b).

4. Simulation Verification of Transferring Pressure Across Goaf Effect

4.1. Model Establishment and Simulation Method

4.1.1. Model Establishment

To analyze the pressure transfer mechanism from working face 1511 mining to BHR 1513, a UDEC numerical model (Figure 4) was constructed based on the mine’s geological conditions. The model spans 337 m horizontally and 120.6 m vertically, with BHR 1513 dimensions defined as 4.0 m (width) × 3.4 m (height). Critical components include the following: strip working face 1513 (60 m width), working face 1511 (200 m width), and dual coal pillars A/B (5 m each). The 3up coal seam exhibits a 3.4 m thickness at a 507 m burial depth, positioned 102 m below the model’s upper boundary. A vertical stress gradient of 0.025 MPa/m was implemented, resulting in 10 MPa/m vertical load imposed on the upper boundary. Displacement constraints were applied as follows: vertical fixation at the base, and horizontal fixation at lateral boundaries. Coal seam occurrence characteristics and mechanical parameters (

Table 1. Distribution and mechanical parameters of strata.

Serial Number	Lithology	Block Parameters		Contact Surface Parameters					Remarks
		Density/kg·m−3	Elastic Modulus/GPa	Normal Stiffness/GPa·m−1	Tangential Stiffness/Gpa·m−1	Friction Angle/°	Cohesion/MPa	Tensile Strength/MPa
26	Fine sandstone	2750	11.3	1020	408	28	14.9	3.8	KS7
25	Medium sandstone	2630	11.7	1180	472	27	15.3	5.1
24	Grit stone	2700	10.1	1090	463	26	14.3	4.2	KS6
23	Mudstone	2300	3.5	462	341	26	8.5	1.6
22	Siltstone	2510	9.6	1040	416	25	13.9	4.3
21	Medium sandstone	2630	11.7	1180	472	27	15.3	5.1	KS5
20	Siltstone	2510	9.6	1040	416	25	13.9	4.3
19	Fine sandstone	2750	11.3	1020	408	28	14.9	3.8
18	Sandy mudstone	2380	4.7	580	232	25	9.7	2.8
17	Siltstone	2510	9.6	1040	416	25	13.9	4.3
16	Fine sandstone	2750	11.3	1020	408	28	14.9	3.8	KS4
15	Mudstone	2300	3.5	462	341	26	8.5	1.6
14	Siltstone	2510	9.6	1040	416	25	13.9	4.3
13	Grit stone	2700	10.1	1090	463	26	14.3	4.2	KS3
12	Mudstone	2300	3.5	462	341	26	8.5	1.6
11	Coal	1260	0.4	280	112	10	2.9	0.5
10	Siltstone	2510	9.6	1040	416	25	13.9	4.3	KS2
9	Fine sandstone	2750	11.3	1020	408	28	14.9	3.8
8	Siltstone	2510	9.6	1040	416	25	13.9	4.3
7	Coal	1260	0.4	280	112	10	2.9	0.5
6	Siltstone	2510	9.6	1040	416	25	13.9	4.3
5	Fine sandstone	2750	11.3	1020	408	28	14.9	3.8	KS1
4	Siltstone	2510	9.6	1040	416	25	13.9	4.3
3	3up coal	1260	0.4	280	112	10	2.9	0.5
2	Siltstone	2510	9.6	1040	416	25	13.9	4.3
1	Fine sandstone	2750	11.3	1020	408	28	14.9	3.8

) were incorporated into the model, with simulations executed using the Mohr–Coulomb yield criterion.

In the discrete element model, the mechanical parameters of block and interface jointly determine the mechanical properties of rock mass. Block parameters include density, bulk modulus, and shear modulus. Interface parameters include normal stiffness, tangential stiffness, cohesion, and friction angle [20]. The bulk modulus K and the shear modulus G in the model are determined by the elastic modulus E and Poisson’s ratio v, and the specific conversion relationship can be found in Formulas (1) and (2).

$$K={\displaystyle \frac{E}{3\left(1-2v\right)}}$$

(1)

$$G=={\displaystyle \frac{E}{2\left(1+v\right)}}$$

(2)

The mechanical parameters of the rock mass obtained from the engineering site are obtained by uniaxial compressive and tensile tests in the laboratory. However, the parameters applied in the discrete element UDEC 6.0 software are the parameters of the rock mass, and it is necessary to convert the rock mechanical parameters obtained in the laboratory. The elastic modulus of block is converted to the elastic modulus of the rock mass by Formula (3); the compressive strength of rock is converted to the compressive strength of the rock mass by Formula (4); and the tensile strength of the rock mass is determined by Formula (5). After the mechanical parameters of rock are transformed into the mechanical parameters of the rock mass, the numerical model is verified. After several parameter verifications and adjustments, the mechanical parameters of each rock strata in the model are determined as shown in Table 1.

$${\displaystyle \frac{E_m}{E_{\mathrm{r}}}}={10}^{0.0186RQD-1.91}$$

(3)

$${\displaystyle \frac{{\sigma }_{cm}}{{\sigma }_c}}={\left({\displaystyle \frac{E_m}{E_r}}\right)}^j$$

(4)

$${\sigma }_{tm}=k{\sigma }_{cm}$$

(5)

4.1.2. Simulation Method

The numerical simulation procedure consists of four main steps: the (1) initial stress analysis of the undisturbed rock mass; (2) sequential extraction of the narrow panel 1513 followed by equilibrium computation; (3) development of the BHR 1513 while maintaining a 5 m protective coal pillar; and (4) incremental extraction of working face 1511 in 10 m intervals, preserving a 5 m pillar (Figure 5), to assess its impact on the adjacent BHR 1513 [21]. As working face 1511 advances rightward to the 90 m position, stress accumulation in coal pillar B triggers its failure. Consequently, the simulation terminates once panel 1511 reaches this critical advancement distance.

4.2. Arrangement of Monitoring Points

4.2.1. Arrangement of Monitoring Points for Vertical Displacement of KS in Overlying Strata

• For the comprehensive assessment of overburden behavior adjacent to BHR 1513, seven vertical displacement monitoring lines (designated as Line A) were installed—one for each key stratum (KS1–KS7). These measurement arrays spanned from the intact coal rib of BHR 1513 to the boundary of Coal Pillar A, with spatial distribution illustrated in Figure 6a (using KS1 as representative case);
• Simultaneously, a dedicated 10 m monitoring line (Line B) was deployed in the immediate roof strata above Coal Pillar B to capture its deformation characteristics during panel 1511 extraction. This instrumentation extended from the pillar’s left edge to a position 5 m inward from its right boundary (Figure 6b).

4.2.2. Arrangement of Monitoring Points for Vertical Displacement and Vertical Stress Above Coal Pillar

To assess the load-bearing behavior of roof strata above coal pillars A and B during the extraction of working face 1511, two monitoring arrays were installed:

Five measurement stations (M1–M5) positioned 2 m above coal pillar A (Figure 6c) to record vertical stress and displacement variations;

An identical configuration (M6–M10) installed at the same elevation above coal pillar B (Figure 6d) for comparative analysis.

4.2.3. Arrangement of Monitoring Points for Surface Displacement of BHR 1513

After the strip working face 1513 becomes a goaf upon the completion of mining, to monitor the change law of surface displacement of the BHR 1513 during the mining of the working face 1511, a monitoring point is arranged on the roof and two ribs of the roadway, D1, D2, and D3, respectively (Figure 6c).

5. Results and Discussion

5.1. Results

5.1.1. Vertical Displacement of KS in Overlying Strata

(1)

The collapse and displacement of 7 KS in the overlying strata of the strip goaf 1513 lateral to the BHR 1513 are shown in Figure 7 (data from monitoring line A), and the following can be learned:

• Before mining the working face 1511, the vertical displacement in monitoring line A of all 7 KS gradually increases to different extents from left to right (extend from the solid coal sidewall of BHR 1513 towards the strip goaf 1513). Among them, the monitoring line vertical displacement of KS1 increases from 279 mm to 374.5 mm, an increase rate of 34.2%. The monitoring line vertical displacement of KS2, KS3, KS4, KS5, KS6, and KS7 sees an increase of 15.9%, 6.8%, 4.8%, 2.7%, 1.0%, and 0.3%, respectively. This increase in vertical displacement is defined as the KS rotation degree of the later Strip goaf 1513 of the BHR 1513;
• When the mining of the working face 1511 is gradually progressed towards the right to 30 m, the vertical displacement in the monitoring line of all 7 KS gradually increases. Among them, the rotation degree of KS1 increases from 34.2% to 35.1%; the rotation degree of KS2 increases from 15.9% to 16.9%; the rotation degree of KS3 increases from 6.8% to 7.9%; the rotation degree of KS4 increases from 4.8% to 5.9%; the rotation degree of KS5 increases from 15.9% to 16.9%; the rotation degree of KS6 increases from 1.0% to 2.7%; and the rotation degree of KS7 increases from 0.3% to 2.4%. It can be concluded that when working face 1511 mines to 30 m, all the 7 KS subside and rotate at the same time with a small rotating increase rate, showing the trend of overall subsidence;
• When the mining of the working face 1511 is gradually progressed towards the right to 60–90 m, the vertical displacement in the monitoring line of all 7 KS increase more obviously. As shown in Figure 7, when the working face 1511 mines to 60 m, the rotation degree of the 7 KS begins to accelerate increasingly, in other words, 60 m is the inflection point of rotation degree of KS.

Figure 7. Vertical displacement of each KS in monitoring line A: (a) KS1; (b) KS2; (c) KS3; (d) KS4; (e) KS5; (f) KS6; and (g) KS7.

Figure 7. Vertical displacement of each KS in monitoring line A: (a) KS1; (b) KS2; (c) KS3; (d) KS4; (e) KS5; (f) KS6; and (g) KS7.

(2)

The collapse of coal pillar B and main roof above lateral goaf 1511 (data from Monitoring line B) are shown in Figure 8, and the following can be learned:

• Before the mining of working face 1511, coal pillar B and goaf 1511 do not exist at this time, and the vertical displacement of rock strata in the measuring line segment KS1 shows a trend of increasing from left to right (incline to strip goaf 1513);
• When the working face 1511 is gradually mined towards the right to 30~60 m, the vertical displacement of rock strata in KS1 continues to increase, and the curve is relatively flat, showing a trend of overall subsidence;
• When the working face 1511 is gradually mined towards the right to 90 m, the vertical displacement of rock strata in the measuring line segment KS1 further increases, showing a trend of increasing from left to right (rotate towards the direction of goaf 1511).

Figure 8. Vertical displacement of KS1 in monitoring line B.

Figure 8. Vertical displacement of KS1 in monitoring line B.

5.1.2. Displacement and Stress of Roof Above Coal Pillar

Before and during the mining of the working face 1511, the change law of vertical displacement and vertical stress of monitoring points M1, M2, M3, M4, and M5 at the 2 m horizons above the roof of coal pillar A is shown in Figure 9; the change law of vertical displacement and vertical stress of monitoring points M6, M7, M8, M9, and M10 at the 2 m horizons above the roof of coal pillar B is shown in Figure 10.

(1)

During the gradual mining of working face 1511 towards the right, the vertical stress of the 5 measuring points above coal pillar A shows an increasing trend, while the increase rate shows a decreasing trend as a whole. When mining to 60~80 m, the vertical stress change curve tended to be flat. As the roof above coal pillar A suffers more pressure, the vertical displacement also increases continuously. Before mining to 60 m, the vertical displacement of 5 monitoring points above coal pillar A shows an increasing trend with the same amplitude. When mining to 60~90 m, the vertical displacement of 5 monitoring points begins to change, showing an increasing trend from M1 to M5, that is, the roof above coal pillar A begins rotating towards the right (rotate towards Strip goaf 1513). Moreover, when the working face 1511 mines towards the right to 90 m, the vertical displacement of the five measuring points differs largely, that is, the roof above pillar A continues rotating towards Strip goaf 1513;

(2)

During the gradual mining of working face 1511 towards the right, the vertical stress of the 5 measuring points above coal pillar B shows an increasing trend, and so does the increase rate. The inflection point arrives when the working face 1511 mines to the distance of 60 m. Similar to coal pillar A, when the roof above coal pillar B suffers more pressure, the vertical displacement also increases continuously. According to the curve variation in Figure 10, when the working face 1511 mines to 90 m, the displacement of the measuring point M6 to M10 (the roof above the coal pillar B from left to right) increases slightly in turn, and the vertical displacement difference of the five measuring points increases, that is, the roof above the coal pillar B continues to rotate towards goaf 1511.

5.1.3. Surface Displacement of BHR 1513

The change law of the surface displacement of monitoring points D1, D2, and D3 of the BHR 1513 before and during the mining of the working face 1511 is shown in Figure 11.

(1)

There is an increase in the displacement of the roof and two coal sides during the mining of the working face 1511 towards the right. The solid coal sidewall sees the largest displacement increase (from 658 mm to 876 mm, an increase of 218 mm), followed by the roof (from 182 mm to 391 mm, an increase of 209 mm) and solid coal sidewall (from 467 mm to 490 mm, an increase of 23 mm);

(2)

When the mining distance of working face 1511 is within 60 m, the roof displacement shows a linear growth trend with a rapid growth rate, while the displacement of coal pillar sidewall grows slowly. When the mining distance of working face 1511 exceeds 60 m, the roof displacement velocity slows down, without obvious displacement, while the displacement velocity of the coal pillar sidewall increases significantly.

5.2. Discussion

Before the mining of working face 1511, in the overlying strata of strip goaf 1513, only KS1 breaks and rotates, KS2 shows a certain degree of deflection subsidence (Figure 7a,b and Figure 12a), and other KS only show a small degree of deflection subsidence (Figure 7c–g). Obviously, the rotation degree of the KS is different (decrease from KS1 to KS7). When the working face 1511 is gradually mined towards the right to 30 m, the KS of overlying strata does not break or rotate due to the small mining length. However, during this process, the overlying strata stress of goaf 1511 transfers to coal pillar B and coal rib along the direction of working face 1511 (Figure 12b). At this time, the pressure of coal pillar B will increase and the overall subsidence will occur (Figure 13b). The overall subsidence of coal pillar B will result in the overall subsidence of the rock strata in the “T” type area, which further leads to the overall subsidence of KS1 and other KS in the strip goaf 1513 (Figure 12b).

When the working face 1511 is gradually mined towards the right to 60–90 m, the compression and deformation of coal pillar B occur due to continuous pressure (Figure 13c,d), resulting in the penetration trend between strip goaf 1513 and goaf 1511, which forms a larger goaf. The overlying strata of goaf break up and collapse in turn, resulting in the breaking and rotation of the original unbroken high KS above strip goaf 1513, which forms dynamic block A and block B (Figure 12c). The broken block A and block B in KS1 above coal pillar A continued to rotate towards strip goaf 1513 under the breaking and rotating effect of high KS (Figure 14), making each KS in the lateral roof of BHR1513 begin to increase the rotation degree (Figure 7). Therefore, the vertical displacement of the 5 measuring points in the roof above coal pillar A begins to change (Figure 9a), showing a trend of rotation towards the strip goaf 1513. The further rotation and collapse of the broken block A and B in KS1 above coal pillar A towards the strip goaf 1513 will exert a disturbance effect on BHR 1513. The law of roadway surface displacement also reveals that when the mining distance of working face 1511 exceeds 60 m, the displacement velocity of the coal pillar sidewall begins to significantly accelerate (Figure 11), which is due to the overall displacement of the coal pillar sidewall towards the roadway space under the rotation of the broken rock in KS1 (Figure 14).

When the mining distance of working face 1511 arrives at 90 m, coal pillar B is eventually crushed (Figure 13d), and goaf 1511 and strip goaf 1513 connect to form a larger goaf. At this time, all 7 KS are broken and rotated with an increased rotation degree, causing stronger disturbance to BHR 1513.

It is worth noting that before the 1511 working face is mined, the KS1 lateral broken rock above coal pillar B is inclined towards the strip goaf 1513. Therefore, when working face 1511 is mined to 30~60 m, the measuring line segment of KS1 gradually changes from an incline towards strip goaf 1513 to rotate towards strip goaf 1511. Therefore, the curve in Figure 8 changes from an inclination towards the left to flat, showing a trend of overall subsidence. When working face 1511 is mined to 90 m, the measuring line segment of KS1 is further rotated and inclined to goaf 1511, so the curve in Figure 8 (when the working face 1511 is mined to 90 m) shows an increasing trend of vertical displacement from left to right (rotated to the goaf 1511).

6. Conclusions

Based on the typical engineering case of gob-side roadway driving with a small coal pillar in the distant face mining, through theoretical analysis and numerical simulation, this paper studied the transferring pressure across the goaf effect and occurrence mechanism of a small-coal-pillar roadway in the distant face mining. This study provides actionable solutions for gob-side roadway stability in distant mining scenarios, directly supporting China’s strategic priorities of ‘safe production’ and ‘green transition’ in the coal sector. The following conclusions are obtained:

(1)

Through theoretical analysis and numerical simulation, this study reveals the “transferring pressure across goaf” effect and its occurrence mechanism in gob-side roadways with small coal pillars under distant face mining. The results indicate that the island coal pillar collapses under the “T”-shaped strata stress after adjacent working face mining, leading to goaf interconnection, the subsequent breaking and rotation of key strata (KS), and severe disturbance to the roadway surrounding rock. These findings align with the research objectives stated in the Introduction, providing theoretical support for stability control in deep gob-side roadways;

(2)

Study Limitations: The numerical model is based on idealized geological conditions, ignoring the impact of complex faults or heterogeneous rock layers. Additionally, potential errors may arise from the laboratory-to-field conversion of coal-rock mechanical parameters, which could affect the accuracy of stress distribution predictions. Future studies should validate the model’s universality by integrating field monitoring data and addressing multi-scale geological complexities;

(3)

Future Research Directions: Subsequent investigations should focus on the dynamic response of roadways under multi-working-face collaborative mining and coupled stress fields. Further optimization of active pressure relief techniques (e.g., directional hydraulic fracturing) and synergy control strategies (e.g., graded anchorage systems) is critical to enhance anti-disturbance capabilities in complex stress environments. Additionally, real-time monitoring technologies and multi-physics coupling models should be developed to quantify spatiotemporal correlations between mining-induced stress and roadway deformation.