Research on the movement law of overlying strata and mine pressure behavior in fully mechanized top-coal caving mining under goaf

During the mining of the lower coal seam under goaf, strong ground pressure behavior and severe roadway deformation occur, affecting normal mining. This study integrated theoretical analysis, laboratory tests, numerical simulation, and field application to explore overlying strata structure/stability, movement and mine pressure behavior laws, and ground pressure sudden change countermeasures in fully mechanized top-coal caving mining under goaf. Results showed that after upper coal seam mining, a stable “masonry beam” formed in the upper key stratum, with floor stress disturbance depth ∼60 m and fractured zone depth ∼7.47 m. Lower coal seam mining initially showed layered overlying strata caving; goaf connection activated upper goaf overlying strata, destabilized the “pressure arch”, and caused shear failure with a “stepped rock beam” structure. Taking goaf connection as the boundary, ground pressure exhibited strong/weak periodic weighting. Weak weighting (larger interval, lower intensity) resulted from lower key stratum periodic fracture; strong weighting (smaller interval, higher intensity) occurred when two coal seams’ damaged areas connected, transferring overlying strata weight to the lower face. When the seam height ratio <5.70 (mining height >8 m), damaged areas connected, causing intense strong weighting. A ratio of 5.70–11.40 (mining height 5–8 m) led to slight weighting intensity difference. A ratio >11.40 (mining height <5 m) prevented connection and ground pressure sudden change. The targeted control scheme achieved good effects, with no support crushing and improved roadway convergence.

1 Introduction

In multi-seam mining areas, when mining the lower coal seam, there appear obvious mine pressure behaviors different from the initial mining, as well as severe roadway deformation and damage, which seriously affect the normal mining of the lower coal seam. In particular, issues such as the distribution characteristics of the mining-induced stress field, the characteristics of overlying strata movement, and the laws of mine pressure behavior of the lower coal seam under repeated mining disturbances have attracted much attention (Song et al., 2021; Yang and Zhao, 2024; Yang and Wang, 2018; Cheng et al., 2019; Chai et al., 2025; Cui et al., 2020; Cao et al., 2025; Huang et al., 2017). Compared with single coal seam mining, due to the change of the original stress in the floor of the coal seam caused by the mining of the upper coal seam, many new phenomena of sudden changes in mine pressure will occur during the mining of the lower coal seam. When the mining-induced stress generated on the floor during the mining of the upper coal seam is superimposed with the mining-induced stress generated on the roof during the mining of the lower coal seam, a superimposed stress field is formed. This exacerbates the damage degree of the overlying rock in the lower coal seam, thereby affecting the safe and efficient mining of the coal seam group (Zhang et al., 2024; Li L. et al., 2023).

Repeated mining of close coal seam will produce secondary disturbance to overlying rock, aggravate the failure of rock stratum, and cause the problems such as ground falling, cracking, and increasing the development height of fracture zones, lead to the destruction of surface vegetation and buildings, and also aggravate the difficulty of rock layer control (Sui et al., 2019; Xie, 2019; Huang et al., 2020; Zhang et al., 2019; Li S. G. et al., 2020). Therefore, it is of great practical significance to study the mechanical behavior and migration evolution of rock formation under the condition of repeated mining of multiple coal seams, and to reveal the failure morphology and mine pressure behavior of overlying strata during shallow buried close multi-coal seam mining, so as to prevent the catastrophe of overburden instability and formulate effective control strategies.

Qian M.G, and Huang Qingxiang et al. (Wang et al., 2025; Huang et al., 2021) studied the transmission law and strata control of dynamic load of rock strata. Yang et al. (2022) studied the evolution law of overlying strata structure in close distance coal seams by similar experimental method. Yang et al. (2019) studied the fracture development law of overlying rock in Gaojialiang Coal Mine by theoretical analysis and experimental method. Li et al. (2020a) used numerical simulation methods to study the overburden rock movement and fracture law under the condition of multi-coal seam mining. Guo et al. (2020) studied the failure characteristics of overlying rock caused by longwall mining by discrete element method. Yan et al. (2022) studied the law of surface movement caused by multi-coal seam mining based on UDEC discrete element simulation software. Yu et al. (2016) studied the rules for roadway deformation during ultra-close coal seam mining.

Yang et al. (2024) Based on a case study of close-multiple coal seams with repeated mining in the Qianjiaying coal mine, a stress distribution model of the floor in the coal seam striking range was established by optimizing the load form of the abutment pressure. Wang et al. (2024) This study utilizes ground-penetrating radar (GPR) to analyze the stability and plastic zone extent of residual coal pillars after upper seam extraction. A theoretical model for the stress distribution of multiple goaves and coal pillars was established. Based on a case study with variable inter-seam spacing, the stress distribution and transfer mechanisms of the floor were analyzed. Yang et al. (2023) physical modelling findings advanced understanding on structural characteristics and stress evolutions of overlying strata over multiple coal seam mining and offered guidance for prediction and mitigation of strata movement associated disasters in underground coal mining with geomechanical and mining conditions similar to those of Buertai coal mine. Li et al. (2020b) by using the ZTR12 geological penetration radar (GPR) survey combined with borehole observations, the overburden caving due to mining of the five coals seams was measured. The research results show that the roof structure formed in the gob area can support the key overlying strata, which is beneficial to ensure the integrity and stability of the upper coal seams in multiple-seam mining of close coal seams. Li et al. (2020d) In order to assess the rationality of the rated shield support capacity (RSSC) experienced selection and guide the reasonable RSSC selection for the subsequent working faces of each coal seam, the coupling relationship between shield and roof strata was revealed during each coal seams mining. According to whether the fractured rock blocks generated by the main roof are articulated and whether the upper coal seam has been mined and influenced on the lower coal seam, two roof structure mechanical models of the rock blocks generated by the thick main roof and two calculation methods of a given load on the rock blocks are proposed. Xiong et al. (2022) investigated the failure mechanism of the lower coal seam working face under the influence of repeated mining in close-distance coal seam groups, with the 17101 working face as the research background. The results show that: affected by repeated mining disturbances, the cracks in the coal face are relatively developed, the strength of the coal body is reduced, and the coal face is more prone to failure under the same roof pressure; During the mining of coal seam 17#, the roofs of different layers above the stope form two kinds of “arch” structures and one kind of “voussoir beam” structure, and there are three different degrees of frequent roof pressure phenomenon, which is easy to cause coal face failure; Under repeated mining of close coal seams, the roof pressure acting on the coal face is not large. The main controlling factor of coal face failure is the strength of the coal body, and the form of coal face failure is mostly the shear failure of soft coal. Kang Z., et al. (Kang et al., 2024) compiles extensive measured data from various mining areas in China to analyze the coupling relationship between the WCFZ development height and six influencing factors: mining thickness, mining depth, coal seam spacing, hard rock lithology ratio, and the slope length of working face. Using principal component analysis and fuzzy comprehensive evaluation, we derived the combined weights of these factors. We developed a prediction model based on multiple regression analysis that incorporates the weighted influences of these factors, further refined into a multivariate nonlinear regression model for greater accuracy. It was successfully applied and validated on the 100,501 working face of Nanyaotou Coal Mine in Shanxi Province, offering a new scientific approach for predicting the WCFZ. Hu and Yu (2024) reveal the failure morphology and mine pressure behavior of overlying strata during shallow buried close multi-coal seam mining, this research takes the Shenmu Zhangjiamao mining area as the engineering background, the study of overlying rock movement and mine pressure distribution law in shallow and close coal seams was carried out. The research results can provide practical experience for similar coal seam group mining in other mines, and can also be used to study the mining pressure behavior of inclined close distance multiple coal seams. Ji et al. (2022) taked a mine in Ningdong as the engineering background, using the research methods of theoretical analysis, numerical simulated, and field measurement, four overburden structure models of single or multiple key strata with repeated mining and single or multiple key strata without repeated mining were constructed. The evolution pattern of bed separation zone was studied from four aspects: bed separation zone development mechanism, formation position, development form, and change of clearance.

Scholars at home and abroad have conducted extensive research on the law of strata behavior in multi-seam mining and achieved abundant research results. However, the research on the mechanism of abrupt strata behavior and corresponding countermeasures in fully mechanized top-coal caving (FMTC) faces under multi-seam mining conditions is relatively insufficient. During the mining process in multi-seam mining areas in central and western China, the problems of abrupt strata behavior such as support crushing and roof collapse occurring in the lower working face lack theoretical basis and effective control measures. Based on this, this paper comprehensively adopts experimental methods such as theoretical analysis, laboratory tests, numerical simulation and field application. Taking the mining conditions of a mine in ordos as the basis, it systematically carries out research in three aspects: the study on overlying rock structure and stability during two-seam mining, the study on overlying rock movement and the law of mine pressure behavior, and the study on countermeasures against mine pressure mutations in fully mechanized top-coal caving mining under goafs. It aims to reveal the overlying rock movement characteristics and the law of mine pressure behavior in fully mechanized top-coal caving working faces under goafs. The expected research results will provide a theoretical basis and reference for control measures for the safe and efficient mining of similar mines.

2 Study on overlying strata structure and its stability

2.1 Analysis of mining-induced influence on the upper coal seam

After the mining of the upper coal seam, the original rock stress in the floor (i.e., the roof of the lower coal seam) changes. Under the action of abutment pressure, the floor will undergo expansion or compression. Therefore, the distribution and transfer law of abutment pressure determine the degree of floor damage, that is, the occurrence state of the roof of the lower coal seam. Abutment pressure acting on the coal seam floor is prone to stress concentration near coal pillars, goafs, coal walls, and open-off cuts. Since coal pillars or coal walls bear the weight of the uncollapsed rock strata above the goaf, the abutment pressure borne by coal pillars or coal walls is relatively large, and their influence on the floor is also more significant in both degree and scope.

According to the theory of elasticity, the distributed force p acting on the plane of a semi-infinite body will have an influence on any arbitrary point M(x,y) below it. The stress condition at any arbitrary point on the floor is shown in Figure 1.

Figure 1

Figure 1. Schematic diagram of force conditions at floor points.

By superposing the distributed force p using calculus, the expressions for the stresses in various directions at any point M under the coal pillar subjected to abutment pressure in the Cartesian coordinate system, i.e., Equations 1–3, can be derived (Yang and Wang, 2018).

${\sigma }_x=\frac{p}{\pi }\times \left(-\frac{y\left(x+a\right)}{y^2+{\left(x+a\right)}^2}+\frac{y\left(x-a\right)}{y^2+{\left(x-a\right)}^2}+{\tan }^{-1}\frac{x+a}{y}-{\tan }^{-1}\frac{x-a}{y}\right)(1)$

${\sigma }_z=\frac{p}{\pi }\times \left(\frac{y\left(x+a\right)}{y^2+{\left(x+a\right)}^2}-\frac{y\left(x-a\right)}{y^2+{\left(x-a\right)}^2}+{\tan }^{-1}\frac{x+a}{y}-{\tan }^{-1}\frac{x-a}{y}\right)(2)$

${\tau }_{xz}=\frac{p}{\pi }\times \left(\frac{y^2}{y^2+{\left(x+a\right)}^2}-\frac{y^2}{y^2+{\left(x-a\right)}^2}\right)(3)$

σ_x—horizontal stress, σ_z—vertical stress, τ_xz—shear stress. x and y are M coordinates; a is the width of the concentrated force application. p—Uniformly distributed support pressure on the coal pillar, in megapascals (MPa); field measurement shows that p = 25 MPa in the 12402 working face. x,y—Coordinates of calculation point M relative to the action boundary of the uniformly distributed force, in meters (m).; a—Coal pillar width, in meters (m).

It can be seen from the above formula that the vertical stress σ_z is the stress with the largest influence range. If the points with equal σ_z in the floor are connected, an “elliptical” contour distribution is formed, and the vertical stress contour of the floor is shown in Figure 2. At 3a below the floor, the vertical stress has decreased to 0.2 times, so the action depth H1 of the abutment pressure in the floor is generally considered to be as per the formula. When y = 3a σ_z = 0.2σ₀ (where σ₀ is the in-situ stress), which conforms to the empirical criterion in mining engineering that “the influence depth of support pressure is three times the coal pillar width” [34]. It is applicable to medium-hard floors with a uniaxial compressive strength (UCS) of 15–30 MPa. In this project, the compressive strength of the floor sandy mudstone is 22.88 MPa, which meets the applicable conditions. Thus, Equation 4 is derived.

$H_1=3a(4)$

Figure 2

Figure 2. Contour lines of vertical stress in the floor.

The floor will develop a plastic failure zone under the action of abutment pressure, as shown in Figure 3. Starting from the location of the abutment pressure peak at point L, the floor is divided into the active stress zone (Zone I), stress transfer zone (Zone II), and passive stress zone (Zone III). After being subjected to the abutment pressure, the stress is gradually transferred from Zone I to Zone III, and the floor reaches a state of limit equilibrium (Qin et al., 2025; Li H. T. et al., 2020; Wang et al., 2016).

Figure 3

Figure 3. Plastic failure depth of the floor.

The failure depth of the plastic zone under the action of ultimate bearing pressure, denoted as H₂, is given by Equation 5.

$H_2=\frac{L\cos \phi }{2\cos \left(\frac{\pi }{4}+\frac{\phi }{2}\right)}{e}^{\left(\frac{\pi }{4}+\frac{\phi }{2}\right)\tan \phi }(5)$

L—The distance from the peak of the abutment pressure to the front of the working face (i.e., the acting width of the concentrated force); φ—Internal friction angle.

Based on the specific mining conditions of the 12402 working face, the width of the coal pillar (a) is 20 m, L is 8 m, and the internal friction angle of the floor rock stratum in the upper coal seam is approximately 33.7°. Substituting these values into Equation 2, Equation 4, and Equation 5, the following results are obtained: the influence range of the abutment pressure on the floor of the upper coal seam working face (H₁) is about 60 m; the plastic failure depth of the floor (H₂) is approximately 7.47 m; the influence of the upper coal seam mining on the lower coal seam working face is about 5.72 MPa, and the influence on the key stratum of the lower coal seam working face is approximately 10.53 MPa. That is to say, the lower coal seam working face and its roof are within the mining influence range of the upper coal seam working face.

2.2 Analysis of overlying strata movement laws during lower coal seam mining

The overlying rock structure of the lower coal seam working face belongs to the structure of a single hard and thick key stratum that has been mined in the upper coal seam, as shown in Figure 4. The lower coal seam working face adopts fully mechanized top-coal caving mining. The large mining height will cause a relatively large rotation of the broken rock blocks of the key stratum, forming a “cantilever beam” structure. Through the process of “stability - instability - restability”, weak periodic weighting will be formed in the working face. When the lower coal seam working face is mined to a certain position, with the large-scale collapse of the key stratum rock blocks, the plastic zones of the upper and lower coal seams will penetrate. The previously stabilized upper key stratum will undergo secondary fracture and instability, and at the same time, the lower key stratum will fracture, forming strong periodic weighting. The mechanism of sudden mine pressure changes in the fully mechanized top-coal caving working face of the lower coal seam is shown in Figure 5.

Figure 4

Figure 4. Overlying rock structure model of the lower coal seam working face.

Figure 5

Figure 5. Mechanism diagram of sudden change in ground pressure on the working face of the lower coal seam.

The discriminant formula for hard rock strata in the structure of a single hard and thick key stratum that has been mined in the upper coal seam is shown in Equation 6 (Yang et al., 2011).

$E_{n+1}{h}_{n+1}^2{\displaystyle \sum _{i=1}^n}{h}_i{\rho }_{i}g>{\rho }_{n+1}g{\displaystyle \sum _{i=1}^n}{E}_i{h}_i^3(6)$

E_i, h_i, ρ_i represent the elastic modulus, thickness, and density, respectively, of the ith rock stratum. During the mining process of the upper coal seam working face, the weighting interval ranges from 17.7 m to 39.7 m. According to the “S-R″ stability theory, the stability of the “masonry beam” structure formed after the key stratum breaks should satisfy Equations 7, 8, 9 (Yang et al., 2011).

${\theta }_1\ge \arcsin \left[\frac{4}{3}\left(i_1-\tan \phi \right)\right](7)$

$h_1+h_1^{\prime }\le \frac{0.15{\sigma }_c}{{\rho }_{1}g}\left(i_1^2-\frac{3}{2}{i}_1\sin {\theta }_1+{\sin }^2{\theta }_1\right)(8)$

$\sin {\theta }_1=\frac{1}{l_1}\left|M-\sum h_1\left(K_p-1\right)\right|(9)$

θ₁-Rotation angle; h₁-thickness, 9.2 m; l₁-length, 26.7 m; i-fracture degree; i₁ = h₁/l₁; tanφ-friction coefficient, 0.3; h₁'-thickness of overlying rock mass, 10.16 m; σ_c-compressive strength,19.27 MPa; ρ₁-density, 2.588 t/m³; M-mining thickness, 4.35 m; ∑h₁-immediate roof thickness, 10.30 m; K_p-bulking factor, 1.38.

By substituting the above parameters into Equations 7, 8, it can be concluded that after the mining of the upper coal seam, the key stratum of the overlying rock forms a stable “masonry beam” structure, and the characteristics of the overlying rock structure after the mining of the upper working face are shown in Figure 6. As the lower working face advances, the failure height of the overlying rock increases gradually, and the failure range of the intermediate rock stratum develops upward in a stepped manner. The migration law of the overlying rock is presented in Figure 7. When the lower working face advances to a certain distance, the intermediate rock stratum is completely fractured, resulting in the connection of the failure zones of the two working faces, as illustrated in Figure 8.

Figure 6

Figure 6. Structural characteristics of the overlying strata after mining in the upper coal face.

Figure 7

Figure 7. Migration law of the overlying strata in the lower fully mechanized caving face.

Figure 8

Figure 8. Schematic diagram of the overlying strata penetration.

3 Study on the laws of overlying strata movement and mine pressure behavior

3.1 Similarity simulation experimental research

3.1.1 Model establishment

The model dimensions of the similar simulation test are 2.5 m × 0.2 m × 1.1 m (length × thickness × height). The actual strike length of the 12402 working face is 200 m. Considering both the feasibility of model fabrication and engineering representativeness, the geometric ratio is set as 1:100. Meanwhile, derived from the similarity principle where ${\alpha }_{\sigma }={\alpha }_L·{\alpha }_{\gamma }$ (${\alpha }_{\gamma }$ denotes the unit weight ratio), the unit weight of the model rock mass is 16.5 kN/m³, and that of the prototype rock mass is 25.7 kN/m³, resulting in ${\alpha }_{\gamma }$ = 1:1.56. Thus, ${\alpha }_{\sigma }$ = 1:100 × 1.56 = 1:156. Derived from the similarity relationship of rock creep ${\alpha }_t=\sqrt{{\alpha }_L}$, $\sqrt{100}$ = 10, meaning 1 day in the model corresponds to 10 days in reality. This matches the on-site daily advance rate of 3–5 m, with the model advancing 3–5 cm per day. The test parameters and scale selections are shown in Table 1. The mixing ratio and thickness of each layer in the model are presented in Table 2. The model uses sand, gypsum, and calcium carbonate as the main materials, and mica sheets are used to simulate the parting surfaces between rock strata. The stratigraphic relationship of the coal seams is shown in Figure 9. The average buried depth of the upper coal seam working face is 199.5 m. Therefore, there is still a 149.84-m-thick rock stratum above the model, and a compensating force of 0.029 MPa needs to be applied above the test bench.

Table 1

Table 1. Similarity parameter.

Table 2

Table 2. Mix proportions of similar materials.

Figure 9

Figure 9. Schematic diagram of the positional relationship of coal seam groups.

To prevent the influence of boundary effects, both coal seams were excavated starting from 40 cm away from the model boundary, with an excavation range of 170 cm for each. In line with the actual on-site situation, the upper coal seam was excavated first. After the movement of the overlying strata tended to stabilize, the lower coal seam was then excavated. Stress measurement lines were arranged at the floor of the upper coal seam and the key stratum between the two coal seams. The specific positions of the measurement lines are shown in Figure 10. The leftmost measuring point was 50 cm away from the model boundary, and the spacing between each measuring point was 30 cm, with a total of 12 measuring points arranged. Stress measurement line 1 was used to monitor the stress changes in the floor during the excavation of the upper coal seam. Measurement line 2 was used to monitor the stress changes in the key stratum during the excavation of the lower coal seam. For stress monitoring, micro pressure sensors are adopted, with an accuracy of 0.01 MPa and a sampling frequency of 1 sample per 30 min (1 sample/30 min); for displacement monitoring, displacement is calculated via image recognition, with an error of less than 0.1 mm.

Figure 10

Figure 10. Schematic diagram of measuring line arrangement.

3.1.2 Analysis of similarity simulation results

As shown in Figures 11, 12, after the mining of the upper coal seam, the maximum subsidence of the overlying strata can reach approximately 3 m. Affected by the lithology of each stratum, the displacement curves of the three survey lines exhibit irregular changes without obvious linear variation or symmetry, but the changing trends are basically consistent. The subsidence of the overlying strata in the middle of the goaf in the upper coal seam is significantly greater than that at both ends, and the overlying strata are damaged in an arch shape. Overall, the overlying strata show an integral subsidence phenomenon, and there are significant differences in the subsidence amounts among different strata, indicating that the overlying strata can still maintain good continuity.

Figure 11

Figure 11. Characteristics of overlying rock caving after mining of the upper coal seam.

Figure 12

Figure 12. Subsidence of the overlying strata after the mining of the upper coal seam.

It can be seen from Figure 13 that when the working face advances to about 45 m away from the measuring point, the floor rock strata begin to be compressed and subside. When the working face continues to advance to about 10 m away from the measuring point, the subsidence of the floor rock strata reaches the maximum. Thereafter, with the advancement of the working face, the floor rock strata begin to expand and heave, and after reaching the maximum value, they will gradually return to the initial value.

Figure 13

Figure 13. Displacement variation of the floor after mining the upper coal seam.

As shown in Figure 14, the mining of the upper coal seam has a wide influence range on the floor stress, and the key stratum 2 is also within this stress influence range. When the working face continuously approaches the measuring point, the stress at the measuring point shows a slow increasing trend. After the working face passes the measuring point, the pressure is released, and the floor stress value drops significantly. When the working face is about 20 m past the measuring point, the stress decreases to the minimum value. As the working face continues to advance, the immediate roof collapses and compacts the goaf, and the floor stress at a certain distance behind the working face gradually recovers to a stable state.

Figure 14

Figure 14. Evolution characteristics of floor stress induced by upper coal seam mining.

From the test results, during the excavation of the upper coal seam, the stress changes in the floor rock strata can be divided into three stages: pressure increase before mining, pressure decrease after mining, and pressure stabilization after mining. During the pressure increase stage, the floor rock strata are compressed and subside; during the pressure decrease stage, the floor rock strata begin to expand; and during the pressure stabilization stage after mining, the displacement changes also tend to be stable.

As shown in Figures 15, 16, when the fully mechanized top-coal caving face of the lower coal seam advances to 102 m, with the initial state of the rock strata unaffected by mining activities as the reference, the overlying strata subsidence values of Survey Lines 1, 2, and 3 do not change significantly, which are basically consistent with the subsidence characteristics after the mining of the upper coal seam. That is, in the early stage of lower coal seam mining, the overlying strata in the upper goaf are basically not subjected to secondary disturbance and remain in a stable state. Survey Line 8 is located at the immediate roof of the sixth upper coal seam; as the fully mechanized top-coal caving face advances, the immediate roof caves immediately after mining, and the maximum subsidence value of Survey Line 8 can reach about 12 m, which is significantly higher than those of other survey lines. Survey Line 7 is located 10 m above Survey Line 8; due to the influence of the bulking property of the caved rock, the maximum subsidence value of Survey Line 7 is about 10.5 m, slightly less than that of Survey Line 8. Survey Line 6 is located at the key stratum, which breaks periodically as the fully mechanized top-coal caving face advances, with a maximum subsidence value of approximately 9 m.

Figure 15

Figure 15. Failure of the overlying strata in the lower coal seam (before penetration).

Figure 16

Figure 16. Subsidence of the overlying strata during mining of the lower coal seam (before penetration). (a) Displacement variation result of measuring (b) Displacement variation result of measuring.

As shown in Figures 17, 18, when the fully mechanized top-coal caving face in the lower coal seam advanced to 121 m, the key stratum in the lower coal seam was completely fractured, leading to a secondary collapse of the overlying strata in the upper goaf. The failure zone extended through to the upper boundary of the model, resulting in a complex interconnected state in the goaf. At the moment of interconnection, the stable “pressure arch” structure formed after the mining of the upper coal seam became unstable, the overlying strata in the goaf were “activated”, and the fractures at the failure boundary widened.

Figure 17

Figure 17. Failure of the overlying strata when the working face advanced to 121 m.

Figure 18

Figure 18. Subsidence of the overlying strata when the working face advances to 121 m. As can be seen from (a) Key Stratum 2 is completely fractured, the intermediate rock stratum collapses fully, and the overlying rock in the upper goaf undergoes secondary collapse. Consequently, the subsidence of the overlying rock along Monitoring Lines 1, 2, and 3 changes significantly again, with the maximum displacement exceeding 8 meters. As can be seen from (b) The variation characteristics of displacement along Monitoring Lines 4, 5, 6, 7, and 8 are basically consistent with those observed before the penetration. However, as the advancing distance of the working face increases, the monitoring range of displacement expands.

As shown in Figures 19, 20, when the working face advances to 154 m, the periodic weighting interval of the working face is about 16 m. Compared with the previous weighting conditions, after the breakthrough, the intensity of periodic weighting increases and the interval decreases, which is referred to as intense periodic weighting. At this time, the intermediate rock strata undergo shear-fall failure, and the rock strata in the upper goaf are dislocated, forming a “stepped rock beam” structure.

Figure 19

Figure 19. Overburden failure when the working face advances to 154 m.

Figure 20

Figure 20. Subsidence of overlying rock strata after mining the lower coal seam. As can be seen from (a), with the further advancement of the working face, the collapse of the overlying rock of the upper and lower coal seams occurs synchronously. The maximum displacement of the overlying rock along Monitoring Lines 1, 2, and 3 exceeds 8 meters, which is consistent with the results shown in Figure 18 (a), while the collapse range is expanded. As shown in (b), the variation characteristics of displacement along Monitoring Lines 4, 5, 6, 7, and 8 are basically consistent with those observed before the penetration. However, as the advancing distance of the working face increases, the monitoring range of displacement is extended.

3.2 Numerical simulation experimental research

3.2.1 Model establishment

Numerical simulation research was carried out using FLAC3D. The Mohr-Coulomb constitutive model was selected for the model, which has dimensions of 400 m × 300 m × 110 m (X × Y × Z) and a mining length of 200 m. There is still a 169.84-m-thick rock stratum above the model. According to the stress gradient of 25 kPa/m, a stress boundary condition of 4.2 MPa was applied to the top of the model, and fixed displacement constraints were adopted for the remaining surfaces. Three stress monitoring lines are set in the simulation experiment with a spacing of 5 m. Five displacement monitoring lines are arranged with a spacing of 10 m, and data are recorded once for every 1 m of excavation.

3.2.2 Analysis on the results of mining pressure behavior laws

As shown in Figure 21, significant stress concentration occurs near the coal walls on both sides of the lower coal seam working face. Compared with single-seam mining, multi-seam mining leads to stress superposition at the coal wall of the lower working face. The stress fields of the left coal wall of the lower coal seam working face and the upper coal seam coal wall are superimposed. Since the right coal wall is located below the goaf of the upper coal seam working face, a triple stress superposition occurs among the goaf, the coal wall of the upper coal seam, and the right coal wall of the lower coal seam working face. This results in the stress peak on the right coal wall being higher than that on the left, which is consistent with the abnormal strata pressure behavior observed at the mining site. It is progressively affected by the stress fields from the upper coal seam goaf and coal wall. When the working face advances to approximately 75 m, it starts to be influenced by the upper goaf; at around 125 m, it is affected by both the upper goaf and coal wall; and at about 150 m, it is influenced by the coal wall of the upper working face. No abnormal strata pressure was observed during the initial mining stage on site, but periodic abnormal strata pressure occurred after mining for a certain distance, which confirms that the numerical simulation results are consistent with the field observations. The weighting interval of the lower fully mechanized caving face in the early stage was approximately 20 m (minor periodic weighting), which is consistent with the on-site mining situation and theoretical analysis.

Figure 21

Figure 21. Vertical stress distribution of the lower coal seam working face.

As shown in Figure 22, the maximum principal stress of the lower strata is relatively small, while that of the upper strata is relatively large. This indicates that the distribution of the maximum principal stress in the overlying strata is non-uniform, with stress concentration areas existing. Due to the relatively large mining thickness of the lower coal seam, it is difficult to form an effectively compacted zone after the collapse of the immediate roof. Moreover, it is challenging for Key Stratum 2 to form a stable pressure arch structure in the goaf. Therefore, Key Stratum 2 cannot effectively support the overlying strata. The upper goaf is affected by secondary mining, and the previously stable pressure arch structure has been damaged.

Figure 22

Figure 22. Maximum principal stress nephogram after mining the lower coal seam.

When extracting displacement curves at 5-m intervals from the roof of the lower coal seam, a total of 15 measurement lines are obtained from top to bottom. Among these, Lines 3-4 are located at Key Stratum 1, and Lines 12–14 are located at Key Stratum 2. The cumulative displacement changes in the overlying strata after the extraction of the fully mechanized top-coal caving face under the goaf are shown in Figure 23. Specifically, Lines 1-6 represent the overlying strata of the upper coal seam goaf. During the extraction of the lower coal seam, the overlying strata in the upper goaf were destabilized and subsided again due to secondary disturbance, with a maximum cumulative displacement of approximately 16.3 m in the roof rock stratum.

Figure 23

Figure 23. Cumulative subsidence of overlying strata.

3.3 Comparison of simulation and field data

The results of peak stress of the key stratum, maximum settlement of the overlying strata, interval of weak periodic weighting, and interval of strong periodic weighting from the similarity simulation test, numerical simulation experiment, and field monitoring are compared; the comparison results are shown in Table 3.

Table 3

Table 3. Analysis of comparison results of different tests.

According to the comparison results, the errors between the results of the numerical simulation experiment, the similarity simulation test, and the field measured results are all within 5%, so the test results are considered relatively reliable.

4 Measures for responding to sudden changes in mine pressure

The main reason for the sudden change in mine pressure in the fully mechanized top-coal caving face under the goaf is that the mining of the lower coal seam causes secondary instability of the upper goaf. The large-scale collapse of the overlying strata leads to the sudden change in mine pressure in the lower fully mechanized caving face. For such accidents, measures can generally be taken from two aspects: reducing the roof load and increasing the working resistance of the support. Combined with theoretical calculations and numerical simulation tests, the reasonable working resistance of the support in the lower fully mechanized caving face after the penetration of the failure areas of the two coal seams is calculated, the reasonable mining height of the lower fully mechanized caving face is analyzed, and a comprehensive control scheme for sudden mine pressure changes is formulated.

4.1 Calculation of support working resistance

When a sudden change in mine pressure occurs during fully mechanized top-coal caving mining under a goaf, the overlying strata in the upper goaf undergo secondary instability, resulting in a “stepped rock beam” structure as shown in Figure 24a. The “stepped rock beam” model is derived based on the rock beam bending theory and the block equilibrium principle. Assuming that the rock beam is a rigid body, the contact hinge points satisfy the shear equilibrium condition. The derivation process is as follows. The calculation of the self-weight load of rock blocks is shown in Equation 10. The Rotational Moment Equilibrium Equation is shown in Equation 11. Herein, based on the actual on-site conditions, the value of θ is set to 4°. Dynamic Load Correction: Considering the impact effect of “activated” rock blocks in the overlying goaf, the dynamic coefficient k_d is introduced with a value of 1.2; thus, the final support force formula is shown in Equation 12. At this time, the support load on the lower fully mechanized top-coal caving face can be calculated using Equation 13. According to previous studies, the mechanical model of the “stepped rock beam” structure is shown in Figure 24b, and the supporting force R1 required to maintain the stability of the roof can be calculated using Equation 14 (Huang et al., 2018; Li Y. et al., 2023; Huang et al., 2022).

$W={\rho }_{2}g{h}_2{l}_2{b}_2(10)$

$W·l_2/2=R_1·l_2\cos {\theta }_2(11)$

$R_1=k_d·{\rho }_{2}g{h}_2{l}_{2}b/\left(2\cos \theta \right)(12)$

$P_m=W+b{R}_1(13)$

$R_1\ge \frac{i_2-\sin {\theta }_{2\max }+\sin {\theta }_2-0.5}{i_2-2\sin {\theta }_{2\max }+\sin {\theta }_2}{P}_1(14)$

Figure 24

Figure 24. Overlying Rock Structure Characteristics of the Lower Working Face. (a) “Step-shaped Rock Beam” Structure. (b) “Step-shaped Rock Beam” Structure mechanics model.

P₁, P₂—Load borne by the rock block; R₂—Support reaction force of rock block N; a—Height of the contact surface; θ₂—Rotation angle of rock block M; Q_A, Q_B—Shear force at the contact hinge point of A, B; l₂—Length of the rock block; Pm—Load borne by the support structure; W—Static load of the intermediate rock stratum; bR₁—Dynamic load of the “activated” rock stratum in the upper goaf; b—Width of the support structure; R1—Support force required to maintain the roof structure.

The load on the support during the sudden change of mine pressure can be calculated by Equation 15, in which some parameters can be calculated by Equation 16 and Equation 17.

$P_m=l_{k}b\sum h_2{\rho }_{2}g+b\frac{i_2-\sin {\theta }_{2\max }+\sin {\theta }_2-0.5}{i_2-2\sin {\theta }_{2\max }+\sin {\theta }_2}\left(h_2+i_2\right)i_2{\rho }_2^{\prime }g(15)$

$\sin {\theta }_{2\max }=\frac{m-0.2\sum h_2}{l_2}=\frac{12.1-0.2\times 9}{27}=0.38(16)$

$i_2=\frac{h_2}{l_2}=\frac{23}{27}=0.85(17)$

l_k—Support controlled roof distance; b—Support width; ∑h₂— Immediate roof thickness; h₂—Thickness of “stepped rock beam” structural block; l₂—Length of “stepped rock beam” structural block; ρ₂g—Bulk density of immediate roof; ρ₂’g—Bulk density of caving main roof; θ₂—Rotation angle of rock block.

According to the on-site mining conditions of the fully mechanized top-coal caving face in the lower coal seam, it can be known that l_k = 6.526 m,b = 1.75 m, ∑h₂ = 9 m, h₂ = 23m, l₂ = 27 m, ρ₂g = 25.5 kN/m3, ρ₂’g = 26.2 kN/m3, θ₂ = 4°.Substituting into Equations 12–14, the load on the support after the sudden change in mine pressure is approximately 19,323 kN. The support efficiency (η_s) of the hydraulic support is calculated as 75%. Substituting this into Equation 15, the reasonable working resistance of the hydraulic supports for the fully mechanized top-coal caving face in the lower coal seam is determined to be 25,764 kN per support as shown in Equation 18.

$p=\frac{p_m}{{\eta }_s}=\frac{19323}{75\%}=25765kN(18)$

4.2 Study on the influence of different mining heights on the laws of ground pressure behavior

By changing the mining height of the lower coal seam to alter the ratio of the interlayer distance to the mining height of the lower coal seam (hereinafter referred to as the “height ratio”), this study explores the influence of different height ratios on the strata pressure behavior patterns in the lower coal seam. When mining height is used as the sole judgment index, with interlayer spacings of 30 m and 60 m between the two coal seams respectively, under the same mining height of 8 m, the former is prone to penetration of failure zones, while the latter is relatively stable. However, traditional parameters cannot quantify this “spacing-mining height” synergistic effect. When interlayer spacing is used as the sole index, it fails to distinguish the differences in overlying strata failure caused by mining heights of 5 m and 12 m. A quantitative critical criterion can be established using the interlayer height-mining height ratio (H/M), which can quantitatively classify three types of periodic weighting: weak periodic weighting (H/M > 11.40), transitional periodic weighting (5.70 ≤ H/M ≤ 11.40), and strong periodic weighting (H/M < 5.70).

With a known interlayer distance of 48.77 m between the two coal seams and under the condition that the upper coal seam has been mined and all other conditions remain consistent, six mining scenarios as shown in Table 4 were developed. Monitoring lines were arranged at the positions of the upper and lower key strata, as illustrated in Figure 25. The UDEC software was employed to construct a numerical model for simulation experiments to analyze the effects of different height ratios on the stress and displacement of the key strata between the layers.

Table 4

Table 4. Test schemes.

Figure 25

Figure 25. Schematic diagram of mining model and survey line layout.

As shown in Figures 26, 27, as the mining height of the lower working face decreases, the variation amplitudes of the stress and displacement of the overlying strata gradually decrease. The variation trends of the two key strata are basically the same, but there are significant differences in their numerical values. When the height ratio is less than 5.70, the stress and displacement of Key Stratum 1 vary particularly significantly. At this time, the disturbance range caused by the mining of the lower coal seam is large, the failure zones of the two coal seams are bound to connect, and the strong periodic weighting of the lower fully mechanized top-coal caving face is relatively intense. When the height ratio is between 5.70 and 11.40, the stress and displacement curves of Key Stratum 1 change within a small range. Since the mining height of the lower coal seam is moderate at this time, even if the failure zones of the two coal seams are connected, the overlying strata will not subside significantly, and the intensities of the large and small periodic weightings of the lower fully mechanized top-coal caving face will not differ greatly. When the height ratio is greater than 11.40, the stress and displacement curves of Key Stratum 1 basically remain stable. At this time, Key Stratum 1 is not significantly disturbed, indicating that a stable supporting structure can be formed in the intermediate rock strata during the mining of the lower coal seam, and the failure zones of the two coal seams will not connect. Therefore, no sudden change in mine pressure will occur during the mining of the lower coal seam. The total mining height of the lower fully mechanized top-coal caving face is 12.1 m, and the height ratio is approximately 4.71 m. As can be seen from the above analysis, the disturbance range during the mining of the lower fully mechanized top-coal caving face is relatively large, resulting in strong periodic weighting, which is consistent with the on-site mining situation.

Figure 26

Figure 26. Comparison of Overburden Stress under Different Mining Schemes. (a) Stress monitoring results of Measuring Line 1. (b) Stress monitoring results of Measuring Line 2.

Figure 27

Figure 27. Comparison Diagram of Overlying Rock Displacements under Different Mining Schemes. (a) Displacement variation results of Measuring Line 1. (b) Displacement variation results of Measuring Line 2.

4.3 Field control effect

4.3.1 Control scheme

In the fully mechanized top-coal caving face of the lower coal seam, the mining-to-caving ratio is 1:2.27, with a coal cutting height of 3.7 m, a top-coal caving height of 8.4 m, and a total mining height of 12.1 m. During the early stage of mining, sudden ground pressure accidents such as roof falls and hydraulic support compressions occurred, and these issues appeared periodically (strong periodic weighting) during subsequent mining operations, seriously affecting the safe production of the working face. To ensure the smooth advancement of the fully mechanized top-coal caving face in the lower coal seam, based on the above theoretical and experimental analyses, a comprehensive control plan for sudden ground pressure changes was developed. The plan mainly includes: adjusting the mining height and mining plan according to the on-site mining conditions; increasing the initial support force of hydraulic supports; utilizing the canopy protection devices of the supports; controlling the advancing speed; and strengthening ground pressure monitoring.

4.3.2 Control effect

The resistance monitoring curve of the hydraulic support in the lower coal seam working face is shown in Figure 28. During the mining period, the convergence values of the roof and floor as well as the two sidewalls of the belt conveyor gateway and the return airway were observed using roof separation indicators. The results of the roadway convergence are presented in Figure 29. According to the on-site monitoring data, the comprehensive control scheme for sudden mine pressure changes has achieved good control effects. No hydraulic support crushing accidents occurred during the mining period. The convergence of the roof and floor of the roadways was less affected by mining activities. The convergence of the roof and floor in the belt conveyor gateway was within 10 mm, and that in the return airway was within 20 mm. The convergence of the two sidewalls in the belt conveyor gateway was within 10 mm, and that in the return airway was within 35 mm. No bolt tray bursting was observed, and no liquid leakage or piston rod shrinkage occurred in the advanced support single props and supports. The comparison table is shown in Table 5.

Figure 28

Figure 28. Monitoring Diagram of Support Resistance. (a) Without control measures. (b) With control measures.

Figure 29

Figure 29. Monitoring curve of surrounding rock convergence in the working face roadway.

Table 5

Table 5. Comparison table of changes before and after control measures.

5 Conclusion

This paper systematically investigates the safe mining issues of fully mechanized top-coal caving faces in thick coal seams under goaf areas. Taking the fully mechanized top-coal caving face in the lower coal seam as an example, the study systematically covers three aspects: the structure and stability of overlying strata in two-seam mining, the movement laws of overlying strata and mine pressure behavior, and countermeasures for sudden mine pressure changes in fully mechanized top-coal caving under goaf areas. The main conclusions are as follows:

1. After the mining of the upper coal seam, a stable “masonry beam” structure can be formed in the overlying key stratum. The depth of stress disturbance in the floor can reach approximately 60 m, and the depth of the floor fracture zone is about 7.47 m. The lower coal seam and the intermediate key stratum remain undamaged but are disturbed to a certain extent by the mining of the upper coal seam. During the initial stage of top-coal caving mining under the goaf, the overlying strata exhibit layered caving characteristics, and the intermediate rock layer is not completely fractured. At this time, the upper goaf remains stable. When the goaves of the two coal seams are connected, the overlying strata in the upper goaf are “activated,” the stable “pressure arch” structure becomes unstable, and large-scale cutting of the intermediate rock layer occurs. During the subsequent mining process of the lower top-coal caving face, the intermediate rock layer is damaged by cutting, the strata in the upper goaf are displaced, and a “stepped rock beam” structure appears.

2. Bounded by whether the goaves of the two coal seams are connected, there are obvious strong and weak periodic weighting phenomena in the mine pressure manifestation. The weak periodic weighting is mainly caused by the periodic fracture of the lower key stratum, with a larger step distance and smaller intensity. The strong periodic weighting occurs after the complete failure of the intermediate rock layer. At this time, the failure areas of the two coal seams are connected, and the weight of the overlying strata in the goaves of both coal seams is transferred to the lower top-coal caving face. The strong periodic weighting has a smaller step distance and greater intensity. When mining the top-coal caving face under the goaf, the pressure arch in the lower goaf develops upward in a stepped manner. When the height of the lower pressure arch reaches the upper coal seam goaf, the stable overlying strata structure in the upper goaf will be disturbed again.

3. The influence of different height ratios on the mining of the lower coal seam was compared by changing the mining height of the lower coal seam. When the height ratio is less than 5.70 (mining height greater than 8 m), the failure areas of the two coal seams are connected. At this time, the disturbance range of the lower coal seam mining is large, and the intensity of the strong periodic weighting on the lower top-coal caving face is relatively high. When the height ratio is between 5.70 and 11.40 (mining height between 5 m and 8 m), the mining height of the lower coal seam is moderate, and the overlying strata will not undergo significant subsidence when the failure areas of the two coal seams are connected. The difference in intensity between the large and small periodic weightings is relatively small. When the height ratio is greater than 11.40 (mining height less than 5 m), the failure areas of the two coal seams will not be connected, so there will be no sudden change in mine pressure during the mining process of the lower coal seam. The mining height and mining plan were optimized based on the on-site mining conditions. After implementing a targeted comprehensive control plan for sudden mine pressure changes on the lower top-coal caving face, good control effects were achieved. No hydraulic support crushing accidents occurred, and the convergence of the roof and floor and the two sidewalls in the belt conveyor gateway and return airway were significantly improved. No bolt tray collapse phenomenon occurred.

Data availability statement

The datasets used and/or analysed during the current study available from the corresponding author on reasonable request.

Author contributions

FD: Writing – review and editing. SW: Writing – original draft, Writing – review and editing. GJ: Writing – review and editing. WZ: Writing – review and editing. HN: Writing – review and editing. LW: Writing – review and editing.

Funding

The authors declare that financial support was received for the research and/or publication of this article. We acknowledge the financial support for this work provided by the Key Laboratory of Xinjiang Coal Resources Green Mining, Ministry of Education (KLXGY-Z2505), the Key Laboratory of Xinjiang Coal Resources Green Mining, Ministry of Education (KLXGY-Z2412), Open Fund of Key Laboratory of Green and Intelligent Mining of Coal Resources in Henan Province (ZCF202508), Doctoral Start-up Fund of Xinjiang Institute of Engineering (2025XGYBQJ07), Xinjiang Uygur Autonomous Region Key Research and Development Program Project (2022B01051-3) and Uygur Autonomous Region Key R&D Program Projects (2024B03017-1).

Acknowledgements

All authors contributed to this article. I would like to thank Co-author of the thesis A. SW, GJ, WZ, LW, HN reviewed and modified the manuscript.

Conflict of interest

Author WZ was employed by Wudong Coal Mine, Xinjiang Energy and Chemical Engineering Co., Ltd., China Energy Group.

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

Generative AI statement

The authors declare that no Generative AI was used in the creation of this manuscript.

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