The stopping line of the working face, also called the terminal mining line, is the position where coal mining stops at the longwall working face. In this position, the support of the working face is about to be withdrawn, and the retracement channel needs to be excavated to evacuate the supports, from which the support area and the retracement channel area together form a large cross-section roadway (Liu et al. 2019; Qin et al. 2021; Gao et al. 2008). Because this large cross-section roadway is produced when the working face stops mining, it is recorded as “a large-section roadway during the stopping period” (LSRSP).

In the lower coal seam mining of close-distance coal seam groups, the mining condition and the overburdened structure of the upper coal seam will affect its mining (Suo et al. 2013; Liu et al. 2020; Wang et al. 2015). The state of the overlying strata of the goaf and residual coal body in the upper coal seam working face is different, which makes the load and periodic pressure of the support during mining in the lower coal seam working face change dramatically (He et al. 2016; Zhang et al. 2018; Cui et al. 2020). This study investigates the selection of a reasonable stopping position of the lower coal seam working face and the support of LSRSP under the interference of a complex residual coal body in the upper coal seam in a close-distance coal seam group.

In the entire life cycle of a fully comprehensive caving working face, stopping and removing supports in a reasonable position is the most important step to end the cycle (Vervoort 2021; Zheng et al. 2019; Lou et al. 2021). Many scholars have conducted extensive and profound research on the many problems of the lower coal seam working face, which is disturbed by the upper coal seam. For coal seam mining under goaf, Pan et al. (2020) believed the roof pressure step distance and advanced support pressure of the working face to be less than those of the working face under the non-goaf. Guo and Yang (2021) used UDEC software to analyze the failure height and collapse characteristics of overlying strata in a thick coal seam under a goaf. Wang et al. (2016) found that the overlying key strata were broken, and the influence range and peak value of abutment pressure decreased significantly only under the influence of the interlayer key strata. Liu et al. (2016) used UDEC simulation to obtain the stress reduction area below the goaf of the upper coal seam, which is less than half of the original stress. For the selection of stopping lines in the working face, Zhu et al. (2021) established a comprehensive evaluation method to determine the reasonable stopping line of the working face by using the stress field excited by the seismic wave. Zhou et al. (2017) determined the structure of a reasonable stopping line of the working face by using a simulation experiment combined with a PAC acoustic emission instrument. Xue et al. (2014) used numerical simulations to determine the optimal stress environment when the stopping line of the working face was 115 m away from the pedestrian roadway. To stop mining at the working face, scholars such as Fei, Li, Lui, and Ma designed to dig a retracement roadway in front of the working face, and analyzed the deformation mechanism and reinforced support of the roadway (Liu and Zhong 2018; Li et al. 2020a, b; Lv 2014; Ma et al. 2018). Li et al. (2020a, b; 2021) excavated two retracement channels in advance to facilitate rapid transportation and support removal; at the same time, a support scheme dealt with the deformation of the surrounding rock of the double channels. Xie et al. (2020) designed and successfully applied the anchored rock beam bearing structure (ARBBS) based on the support requirements of a large-section coal roadway. Meng et al. (2016) designed the technical scheme of “three anchors” combined support, which comprises the primary support of full-section anchor net and shotcrete, and the secondary reinforcement of high pre-stressed anchor cable and anchor grouting, according to the characteristics of the relatively broken surrounding rock. Huang and Gao (2013) adopted a combined support scheme for high-strength bolts, combined anchor cables, W-shaped steel belts, metal meshes, and concrete according to the characteristics of broken top coal in fully mechanized top coal caving mining.

The above-mentioned scholars have contributed to the development of multi-coal seam mining, stopping method of working face, and support of large cross-section coal roadway(Xie et al. 2022a, b; Chen et al. 2022). However, research on the reasonable position of the lower coal seam stopping line disturbed by close upper coal seam goaf and remaining coal body synthesizes many factors. The selection method of stopping lines in a single coal seam is not suitable for multi-coal seam mining (Zhang et al. 2021; Cheng et al. 2020). At the same time, the traditional advanced extracting retracement channel method has some problems with channel support and the risk of rock bursts (Lv 2014; Li et al. 2021). Under the strong influence of the end-mining coal pillar in the overlying close-distance coal seam, this method is abandoned in the tunneling of the retracement channel in the underlying working face.

In summary, based on the geological conditions of close coal seam group mining in a coal mine, the reasonable stopping position of the lower coal seam N316 working face and the support method of LSRSP are studied. The mechanical parameters of coal and rock, the three-dimensional corresponding relationship between the upper and lower coal seams, and the periodic pressure of the working face are determined through on-site observation and laboratory experiments, thus focusing on the core problem of selecting a reasonable position of the lower coal seam stopping line under the interference of the overlying thick coal seam end-mining coal pillar. Through numerical simulation and similar simulation experiments, the stress arch structure and movement of the overlying strata under different stopping positions were obtained, and the evolution process of the stress arch structure of the overlying strata in the underlying coal seam is verified by multiple stress parameters. Therefore, the roof stress state of the retracement channel in the underlying working face is analyzed, and a reasonable stopping position under this working condition is obtained through comprehensive comparison. At the same time, the technology of self-excavating the retracement channel is adopted in the underlying working face, based on the different control requirements of the roof above the support and the roof of the retracement channel in the stopping area, through a comprehensive analysis of the prestress field simulation of the support scheme, the partition asymmetric support scheme of LSRSP is proposed. After on-site implementation, the support was withdrawn safely and smoothly, providing a method for guidance and engineering reference for realizing reasonable stopping in the working face under similar geological conditions.

2.1 Geological conditions

The production capacity of the mine reached 4.8 million tons per year. The development way of the well-field system is a main inclined shaft and auxiliary vertical shaft. Mining No. 4 coal seam of Permian Shanxi Formation and the No. 3 coal seam and the No. 5 coal seam of Carboniferous Taiyuan Formation were at + 1035 m, + 885 m, and + 825 m levels, all with low gas content in the coal seam. the coal mining technology is fully mechanized top coal caving. As shown in Fig. 1, the depth of the N316 working face in the No. 3 coal seam is approximately 400 m, the strike length is 2517 m, the dip length is 180 m, and the dip angle is 2°. The average thickness of the coal seam is 5.30 m, the mechanical mining height is 3.2 m, and the coal caving height is 2.1 m. The upper 25 m is the N416 and N418 goaf of the No. 4 coal seam, which belongs to the mining situation under the goaf of the lower working face of the close coal seam group.

Position relationship of short-distance coal seam and coal-rock columnar chart

2.2 Working condition

2.2.1 Influence of overlying goaf and end-mining coal pillar

As shown in Fig. 2, the N316 working face is now vertically mapped to the overlying coal seam, and it was found that 78% of the area is mined out, and the left end-mining coal pillar is in front of it. The different overburden structures of the goaf and end-mining coal pillars in the N416 and N418 faces will have a direct impact on the stopping of the N316 face. From the profile, it can be seen that stopping in different positions of the N316 working face corresponds to different overlying rock states, from which three types of stopping areas are obtained:

• (1) Area I: stop mining under the mined-out of N416 and N418 and staggering the upper stopping line externally, referred to as ESUL.

• (2) Area II: stop mining in the area below the stopping line of N416 and N418, and overlapping with the upper stopping line, referred to as OUL.

• (3) Area III: stop mining under the entity coal of the remaining end-mining coal pillar of the No. 4 coal seam and staggering within the upper stopping line, referred to as ISUL.

Schematic diagram of overlying coal and rock in N316 working face of the No.3 coal seam

2.2.2 Selection requirements for stopping coal caving time

When the working face stops mining, not only should the influence of different stopping positions be considered, but also the coal caving should be stopped timely according to the motion state of the basic roof plate. As shown in Fig. 3, the last pressure before stopping is judged according to the support pressure monitoring, which can be evacuated after the scraper conveyor stops caving coal. Then, the working face continues to advance, and the re-advancing distance is evaluated according to the periodic weighting interval to realize the self-excavation of the retracement channel under a stable basic roof suspension plate.

Schematic diagram of the movement of key blocks in overlying strata when stopping coal caving

2.2.3 Requirements for each stage of the stopping operation

The traditional workface stopping operation consists of two major steps: selection of the stopping position and support of the LSRSP. As shown in Fig. 4, a reasonable stopping operation should have the following six phases and six accompanying requirements.

Schematic diagram of requirements for each stage of stopping operation

• (1) Stage a: It is necessary to preliminarily determine the stopping position where the overlying rock structure is relatively stable according to the coal and rock conditions on the coal seam, and in this study, it is necessary to consider the influence of the stopping line of the overlying No.4 coal seam.

• (2) Stage b: To form a reasonable stopping coal caving distance, it is necessary to determine a reasonable stopping coal caving time according to the movement state of the overlying strata on the working face, stopping coal caving to help support the breaking of the overlying main roof is very important for reducing the support pressure during mining.

• (3) Stage c: With the movement of the hydraulic support, the laying of the metal mesh to support the broken top coal is a key step to ensure the smooth removal of the hydraulic support and control the safe collapse of the roof after the removal of the hydraulic support.

• (4) Stage d: The retracement channel and support area together constitute the LSRSP. It is necessary to effectively support the roof and coal wall of the retracement channel to prevent the fall of ground and spalling rib accidents.

• (5) Stage e: During the hydraulic support removal step by step, prevent large-area roof collapse of the LSRSP from causing hydraulic support pressing and roof accidents, and it is required to follow the hydraulic support removal operation to cover and remove the hydraulic support.

• (6) Stage f: After the support is withdrawn, prevent the roof of the evacuated support area from sinking excessively and crushing the support, thereby making it difficult to remove the next support; it is required to move the support and coal caving in turn until the withdrawal is completed.

Through rock mechanics experiments and field observations, the periodic weighting step, weighting strength, top coal fissure development status, and coal and rock mechanics characteristics of the working face were comprehensively determined.

3.1 Measurement of coal mechanical parameters in N316 working face

To master the mechanical parameters and coal quality of the No. 3 coal seam, a drilling core was carried out in the roadway on both sides of the N316 working face and brought back to the laboratory to make standard rock samples, and density, uniaxial compression, and shear splitting tests were carried out. As shown in Fig. 5, the experimental data were averaged to obtain the parameters shown in Table 1.

Measurement of parameters in laboratory rock mechanics experiment

It is known that the No. 3 coal seam is bituminous coal and the coal quality is weakly sticky coal. From the coal mechanical parameters shown in Table 1, the uniaxial compressive strength of the coal is 14.24 MPa, the cohesion is 2.88 MPa, and the internal friction angle is 34.32°, which indicates that the coal is relatively hard. Considering that the thickness of the top coal without caving is 2.1 m at the stop mining stage, which can be obtained from the mechanical parameters, the hard coal quality makes the top coal still have a certain degree of integrity instead of being completely destroyed into crushed coal. Therefore, in the later support design, because the support needs to be evacuated, the roof should not be too strong to avoid the top coal hanging and can not sink and collapse so as to crush the support that has not been evacuated.

3.2 Borehole observation on surrounding rock condition of the roof of N316 working face

On-site observations show that the periodic weighting interval of the working face is 17–26 m, and the fracture development of the top coal above the support was measured by drilling peeping, as shown in Fig. 6.

Location and peep view of boreholes arranged on site

Through the borehole peep at a depth of 3 m in front of the support, it was found that there were many cracks in the coal body of the roof, the entire top coal was broken, and the rock part was relatively complete. The crushing condition of the top coal is different in different measuring positions, for example, the peep view near the No. 30 hydraulic support shows a large number of broken belts concentrated in the middle of the top coal, and it is speculated that there may be large transverse fractures here; the peep view at the No. 70 support shows that the coal body in the shallow part of the borehole was broken in a large scale, and with many open cracks, the coal block is easy to fall; the peep view at the No. 110 support shows that the top coal was soft, and there were many broken coal and collapse holes at the boundary of roof coal and rock, and many cracks in the rock mass. Comprehensive observations show that the fractured zone of the roof is more than 2 m, and there is a possibility that the entire top coal will fall off in a local area, making it difficult for the bolt support to play an effective role.

To determine the reasonable stop-mining position of the lower coal seam in the close-distance coal seam under the influence of the overlying end-mining coal pillar, the three basic position relationships of stopping in the close-distance coal seam were theoretically analyzed, and the suitable stopping position was determined by numerical simulation, similar simulation, and field practice.

4.1 Interaction of overlying strata at different stopping positions

As shown in Fig. 7, the characteristics of the three basic stopping positions and the interaction of their main roof key blocks in the N316 face are analyzed as follows.

Schematic diagram of interaction relationship between key blocks at different stopping positions

• (1) ESUL: the main roof of the upper coal seam was broken, and the goaf was compacted. When mining the lower coal seam, the main roof of the lower coal seam is broken periodically, which makes the main roof of the upper coal seam prone to secondary breaking, turning, and sinking, and its weight acts on the key blocks of the lower coal seam. When mining is stopped at this time, the support may directly bear the pressure of the basic top key blocks of the upper two layers of coal.

• (2) OUL: the key block of the main roof of the upper coal seam has a secondary rotation and the rotation angle reaches the maximum; part of the force squeezes the broken key block laterally, and the goaf is compacted with the key block of the lower coal seam, and the support does not act as the main body of the load.

• (3) ISUL: the overlying strata fracture structure under this condition is basically the same as that of single-layer coal mining. At this time, the main roof of the upper coal seam is a key layer overlying the working face of the lower coal seam, and its key block generally lags behind the fracture of the key block of the main roof and rotates and extrudes the backside of the key block of the main roof.

4.2 Numerical simulation analysis of different stopping positions

4.2.1 Establishment of numerical model

To analyze the fracture pattern and stress field distribution of the overlying strata at different stopping positions, a discrete element numerical simulation method was used to construct a calculation model with the same size as the actual size in the field, and the hydraulic support model of the N316 working face was constructed at the same rate, as shown in Fig. 8. The model size is 250 m × 88.1 m, the No.4 coal seam is 38 m away from the top of the model, and the studied coal section is 75 m away from the boundary of both sides of the model, so the influence of size effect can be ignored. The buried depth of the No.3 coal seam is about 400 m, so the vertical stress of 8.5 MPa is applied to the upper boundary of the model, and the lateral pressure coefficient is 1.2 to obtain the horizontal gradient stress of 10.2–12.4 MPa. The No. 4 coal seam in the model is first mined to the preset stopping line, and then the lower coal seam was mined. Near the stopping line of the upper coal seam, different fracture line locations of the main roof may affect the simulation results when mining in the lower coal seam. The following three fracture line positions (types I, II, and III) are preset for the simulation to obtain the numerical calculation results of different stopping positions under different fracture positions of the main roof.

Schematic diagram of the numerical model

4.2.2 Formation of stress arch and stress superposition of arch foot

As shown in Fig. 9, the numerical simulation results clearly show the structural and morphological characteristics of the stress arch of overlying strata below three forms of stopping positions, and at the identical time, the stress concentration degree of surrounding rock of the retracement channel below every stopping position is obviously compared. The evolution characteristics of the arch stress at different stopping positions were as follows:

Simulation diagram of stress arch evolution

• (1) When the stopping position is externally staggered at a large distance from the upper stopping line, the overburden of the stopping area has a front and rear double-stress arch structure: the rear stress arch with the rear goaf and the roof of the retracement channel as the arch foot; the front stress arch with the upper stopping line and the roof of the retracement channel as the arch foot. At this time, the stopping position of the underlying coal seam is at the superposition of the front arch foot of the rear arch and the rear arch foot of the front arch; therefore, the pressure on the roof of the retracement channel of the underlying coal seam is the largest, and the stress environment is not conducive to the safety control of the top plate of the LSRSP.

• (2) With the shortening of the distance of external staggering, the front stress arch with the upper stopping line and the roof of the retracement channel as the foot of the arch is gradually smaller, and the back stress arch with the rear goaf and the roof of the retracement channel as the foot of the arch is gradually larger, but the superimposed strength of the foot of the two arches was still stronger, which was still not conducive to the safety control of the top plate of the LSRSP.

• (3) When the stopping position overlaps with the upper stopping line, the front arch dies and the overlying rock in the stopping area has a single arch structure, and the retracement channel of the underlying coal seam is under the protection of a single arch, exerting relatively minimal roof pressure.

• (4) When the stopping position is staggered internally on the upper stopping line, the single-stress arch structure gradually expands before the overlying rock structure on the working face starts to break, and the pressure on the top plate of the retracement channel is relatively small.

• (5) With the gradual increase in the internal staggering distance, the overlying rock in the stopping area is a nested double-stress arch structure consisting of a large arch and a small arch, which takes the rear goaf and the roof of the retracement channel as the arch foot, while the small stress arch takes the upper stopping line and the retracement channel as the arch foot. Although the stress of the double-arch structure is superimposed on the front arch foot, resulting in a larger stress concentration in front of the retracement channel, the pressure on the top plate of the retracement channel is smaller than that of the external staggering.

4.2.3 Stress at each stop position for different prefabricated fracture lines

To show the stress status of different fault lines and stopping positions in detail, three indices of vertical stress, maximum shear stress, and the second invariant deviatoric stress (J₂) were introduced to show the stop stress environment (Zhang et al. 2014; Wang et al. 2021; Chen et al. 2020). According to the numerical calculation results shown in Fig. 10, the following features can be observed:

Stress program under different fault lines and different stopping positions

• (1) From the comparison of the programs of the three stress indexes at the same stopping position but different fault lines, it can be observed that the fault line positions have no apparent influence on the overall stress shape, location, and peak value, but there is only a small difference in the degree of stress concentration(this is due to the difference of interaction between overlying rock blocks caused by different positions of fault lines), which is the same in general.

• (2) Stopping at an external-staggered distance of 50 m from the upper stopping line, the cloud pictures of the three indexes all show the double-stress arch structure and the superimposed arch foot at the top plate of the retracement channel. The vertical stress, shear stress, and J₂ at the top plate of the retracement channel all exhibit the characteristics of high depth, and the longitudinal depth is more than 30 m, where the J₂ peak value belt is tilted to the gob side; therefore, it can be concluded that the continuity and high value of longitudinal stress transfer caused by stopping under the goaf, which is extraordinarily unfavorable for stopping operation.

• (3) Stopping directly below the upper stopping line, the cloud pictures of the three indexes all show a stress arch, and the whole stress environment is optimal; the stress concentration at the top plate of the retracement channel is relatively low, and there is no vertical development, which is suitable for stopping in terms of stress.

• (4) Stopping at an internal-staggered distance of 50 m from the stopping line, the cloud pictures of the three indexes all show a large and small nested double stress arch structure, and the stress superposition of the front arch foot is apparent; the vertical stress concentration area of the coal wall in front of the retracement channel has a large transverse width and the local concentration of shear stress, while J₂ has both characteristics.

4.2.4 The law of displacement evolution in different stopping positions

As shown in Fig. 11, the overlying strata displacement due to underlying seam coal mining is simulated, it is known that there are three distinct displacement zones in the displacement field of the overlying strata in the working face under the three stopping positions: the goaf stability area is the maximum displacement area—the non-caving coal area is the larger displacement area—the retracement channel area is the smaller displacement area. The maximum displacement of the overlying strata caused by the mining of the underlying coal seam is about 6 m, and the regional span of overlying rock displacement zoning of ESUL in the underlying working face is obviously larger than that of ISUL, but the shape is basically the same. However, when it is close to overlapping, the shape of the zone is obviously different, and the longitudinal direction is no longer continuous but dislocated, indicating that the overburden in the upper coal seam acts on the roof of the retracement channel directly below it, which is not conducive to the stability of its roof.

Displacement simulation nephogram of different stopping positions

4.2.5 Analysis of reasonable stopping positions based on numerical simulations

The above analysis shows that the stress environment of the underlying working face is relatively good when stopping mining at the position of OUL and ISUL-SD. An elaborate simulation is performed to explore the reasonable stopping position, as shown in Fig. 12, which is summarized as follows:

Fine simulated stress contrast of the area under the upper stop line

• (1) This area is in the process from external-staggered to overlap and then to internal-staggered, vertical stress, shear stress, and J₂; the front and rear double stress arch → the front arch gradually decreases → the front arch dies, and the double arches are combined into a single arch → the single arch expands the scope.

• (2) Through a comprehensive comparison of the stress environment of the three indices, OUL and ISUL-SD are the reasonable stopping positions of the stress angle, and the stress concentration degree of the roof of the retracement channel is the lowest.

• (3) As shown in Fig. 12, once the internal staggered distance is too large, the stress of the front foot of the stress arch is secondary-concentrated, which is caused by the fracture of the cantilever structure. In addition, an excessively long internal staggered distance is also limited by the width of the end-mining coal pillar.

4.3 Similar simulation experiments of different stopping positions

Based on the results of the site investigation, a 1:120 similarity ratio model was established on a large-scale two-dimensional similarity simulation platform in the laboratory. The model was excavated from left to right, and the left boundary prefabricated the empty side to achieve rotation and sinking of the main roof. The model first excavates the No.4 coal seam and then excavates the No.3 coal seam situated below after reaching the designed stopping line.

As shown in Fig. 13, the following rules were obtained through similar simulation experiments:

• (1) Figure 13a and b show the mining and stopping stages of the overlying No. 4 coal seam. The total non-coal caving distance at the stop mining stage is about the periodic weighting distance (it is also necessary to stop caving to maintain the stability of the roof at the stopping stage of the No. 4 coal seam working face, which is determined by the mining of this working face).

  

  Similar simulation experimental diagram of different stopping positions

• (2) Excavation of the No. 3 coal seam from Fig. 13c shows main roof fracture line in front of the hydraulic support, secondary breakage in the main roof of the upper coal seam, which then slides and sinks with the main roof of the lower coal seam acting on the support.

• (3) Figure 13d shows the retracement channel is externally staggered with the upper stopping line(ESUL). When the lower stopping line is on the external-staggered side, the main roof fracture line is located in front of the support, and the supports are compacted by entire high depth broken rock formation, which is not suitable to remain there for a long time and is removed.

• (4) Figure 13e shows the retracement channel overlaps the upper stopping line(OUL). When the upper stopping line is overlapped, the rock formation behind the support of the working face breaks integrity, and the goaf is compacted, while the supports are just standing below the cantilevered structure and protected by it.

• (5) Figure 13f shows that the retracement channel is internally staggered on the upper stopping line within a short distance(ISUL-SD). When the internal-staggered distance is small, the main roof fracture line is above the support, however, the area without coal caving plays a bearing role in breaking the main roof to a certain extent and slows down the support pressure, and the overburden structure above the solid coal is still stable.

• (6) Figure 13g shows the gradual increase of the internal-staggered distance. With the gradual increase in the internal-staggered distance, the overburden structure above the stope space began to break.

• (7) Figure 13h shows that the retracement channel is internally staggered on the upper stopping line within a long distance (ISUL-LD). The main roof of the upper coal seam started breaking caused by the lower coal seam, which then leads to the fracture of the large structure of the entire high-depth overlying rock. And the fracture lines are all above the support, indicating that the support would not stay for a long time and remove.

4.4 Comprehensive simulation and on-site selection of actual stopping position

Based on the above numerical simulation results and physical similarity simulation experiments, it is concluded that OUL and ISUL-SD are suitable for stopping mining in the stress field and overlying strata structure, which is the reasonable stopping position of coal seam working face at close distance.

The above simulation shows that overlapping and short-distance internal staggering are suitable for stopping in the stress fields and overburden structures. However, according to field experience, when the stopping lines overlap, there is a hidden danger that the key blocks on the working face will slide and cause instability to the entire structure. Therefore, the actual stopping position is selected as the internal-staggered stopping line within a short distance. The detailed selection process of parking and stopping location is shown in Fig. 14. The process flow of stopping mining and stopping coal caving aims to make the stop mining position not only have the optimal surrounding rock stress environment but also reduce the rotation angle of the broken basic roof and bear part of the load of the basic roof under the stable cantilever structure. Among them, when the working face detects the weighting phenomenon of basic roof breaking, it starts to stop caving until the weighting ends. According to the quantitative relationship, the distance of no-coal caving behind the support in the site is 10 m, which ensures the stability of the overlying strata in the stopping stage.

Schematic diagram of reasonable stopping position and non-caving process

The support design of the LSRSP is carried out based on the selection of the actual stopping location. Because the support design should serve the withdrawal process, it is not suitable to use the traditional support idea to design the scheme, and the dynamic process of support evacuation should be considered.

5.1 Support design principle based on support withdrawal technology

A reasonable support design can guarantee the operational requirements of each stage on site. The process flow of the hydraulic support withdrawal process is illustrated in Fig. 15. Through the fine division of the support required by the sequential frame dismantling process in each stage, the key points of support in the three periods are proposed:

Support period divided by the support withdrawal process

• (1) In the early stage, the purpose is to support the LSRSP and prevent the roof spanning 8.68 m from falling partially and to maintain integrity. This emphasis is on the control of overlying broken top coal in the retracement channel, and the support at this stage should combine strength with weakness to avoid excessively strong support of the roof.

• (2) During the withdrawal period, the purpose is to dynamically follow the supporting working space and to ensure the stability of the roof of the dismantling working position and passageway. In particular, because the personnel are concentrated in the area where the racks are being withdrawn, it is necessary to set up shield supports to cover the entire operation and prevent racks from being withdrawn in advance.

• (3) At the later stage, the purpose is to lap the fallen roof coal after the support is withdrawn, in order to prevent the top coal from falling directly and crushing the adjacent support. At the same time, it is necessary to prevent the suspended top coal from falling, so as to prevent the large-area top coal from suddenly falling and causing a strong mine pressure disaster.

5.2 Asymmetric support idea of strong and weak partition of the roof

According to the withdrawal period of approximately 35 days and the requirements of the withdrawal process, the idea of asymmetric support for the strength and weakness partition of the roof of the LSRSP is proposed, and its principle is shown in Fig. 16. The roof of the LSRSP is divided into the support roof area and the retracement channel roof area. The "strong and weak" refers to the different support strength required by the supports and channel roof. The roof of the support should not be too strong, and excessive support will prevent the roof from free collapse after the supports are removed, thereby, preventing the formation of a large area of the suspended roof and serious harm. Meanwhile, there is no support in the retracement channel area, necessitating the strengthening of the roof in this area to avoid roof fall. “Asymmetric” means that the support schemes designed for the roof area of the support and the roof area of the retracement channel are not of the same strength (strength or weakness), and are asymmetric, in order to meet the requirements of roof management of the two areas in the withdrawal operation.

Principle diagram of the partition asymmetric support idea

5.3 Analysis of key factors of support design

Under the guiding ideology of asymmetric support for strong and weak partition of the roof, the key of support is to ensure that the stopping mining position is under the stable cantilever structure of overlying strata. Based on this, the integral roof net structure is constructed, so that the strong and weak partition support of the roof is formed as a whole. As shown in Fig. 17, the foundation of effective support is that the cantilever plate above the stop position has a stable structure. During the entire dismantling process, the roof of the bracket is supported by the early bracket and the later woodpile, and the roof of the passageway is supported by the physical coal gang and the shield bracket during dismantling, both of which require better roof integrity. Therefore, the layout of the double-layer metal mesh to achieve strict watch protection, the key to its implementation is to determine the reasonable stopping coal caving position and lay the mesh in time.

Schematic diagram of support design

5.4 Support scheme and implementation effect

5.4.1 Scheme of support in each period

The partition asymmetric support scheme designed according to the withdrawal process is shown in Fig. 18.

Scheme diagram of time-sharing partition support

• (1) Strengthening support in the early stage: All roofs are supported by anchor cables, and three anchor cables are laid on the roof of the support and the channel. Although it is important to support the roof of the channel, dense support should not be adopted. Therefore, the combined anchor cable is designed to anchor the roof with local strength to achieve a supporting effect on the basis of reducing consumables. To support the middle area of the long-span space, a single hydraulic prop was set up to enhance the roof protection ability. The sidewall of the channel was set with anchor rods to prevent the crushed coal from extruding out. A double-layer metal mesh is laid on the entire section to prevent the roof from falling. The specific parameters selected by each supporting material are shown in Fig. 18.

• (2) Follow-up support during the withdrawal period: To ensure the stability of the roof in the working space for frame removal and protect the safety of the workers, the frames are removed from one end of the support group to the other end in turn, and two shield supports are set to support the roof during the entire process.

• (3) Weakening support in the later stage: After withdrawing the bracket, the wooden pallets were staggered in time to undertake the sunken roof, and then the shelves were removed in turn and continued to be erected. It is worth noting that because the roof above the bracket is anchored by anchor cables if the roof does not sink and hangs after the bracket is removed, it is necessary to disassemble the support for the caving operation to prevent the hanging roof from crushing the bracket to be removed and to avoid a disaster caused by the sudden collapse of the large-area roof.

5.4.2 Support the pre-stress field

As shown in Fig. 19, a three-position prestressed field of the anchor cable and combined anchor cable support was constructed. By comparing the normal anchor cable support section, combined anchor cable support section, and their combined support section, it can be seen that the stress field of the channel roof is stronger and has good continuity, whereas the prestressed field of the roof above the bracket is weaker, thus avoiding strong support and realizing the strong and weak support of the roof position in the LSRSP where mining is stopped. At the same time, the supporting stress of the combined anchor cable is obviously stronger than that of the common anchor cable, and the depth of the supporting high-stress area is approximately 3.5 times that of the common anchor cable, and the range of the high-stress area combined with the common anchor cable is increased by 2 times, which can achieve a better supporting effect by saving the number of anchor cables.

Cloud picture of the prestressed field of large section space support

5.4.3 Field application effect

Figure 20 shows the entire process from the excavation of the retracement channel to the formation of LSRSP and then to the safe and smooth removal of supports. Under the condition of reasonable selection of stopping location and the partition asymmetric support scheme, the N316 working face is controlled within 35 days from support to completion of withdrawal, which saves 10 days. At the same time, the combined anchor cable is used to realize the strong support of the roof of the retracement channel, and the protective effect of double-layer metal mesh on the broken coal of the roof is obvious. Therefore, it is avoided that the support is difficult to be pulled out due to the large deformation of the roof during the withdrawal of the support, the section maintenance and the secondary reinforcement anchor cable support are avoided, thus saving the use and labor cost of the anchor cable drilling machine, and it is estimated that the labor cost and material cost will be saved by CNY 235,000. In addition, according to the monitoring data of the displacement of the roof and side parts of the retracement channel, when the support is completed for about nine days, the deformation tends to be stable, and the maximum subsidence of the roof before removal is controlled at about 140 mm, which does not affect the withdrawal of the support and ensures the speed and efficiency of the withdrawal process. In summary, the support plan designed according to the withdrawal process achieves effective control of the LSRSP roof with less consumables and high safety.

Field application effect and data monitoring

1. (1)

   The structural characteristics of stress arch at different stopping mining line positions are as follows: ① ESUL: the overburden of the stopping area is a front and rear double stress arch structure, and the stopping line of the underlying coal seam is at the superposition of the front arch foot of the rear arch and the rear arch foot of the front arch. ② OUL: the overlying rock in the stopping area has a single arch structure. ③ ISUL-SD: it is also a single-arch structure. ④ ISUL-LD: the overburden of the stopping area is a nested double-stress arch structure consisting of a large arch and a small arch.

2. (2)

   The multi-stress indicators of maximum principal stress, vertical stress, shear stress, and deviatoric stress second invariant are verified to obtain the evolution process of the stress arch structural form of the overburden rock of the underlying coal seam (ESUL → OUL → ISUL-SD → ISUL-LD) as follows: a front and rear double stress arch structure → the front arch gradually decreases → the front arch dies out, and the double arch synthesizes the single arch → the single-arch range expands → the nested double arch.

3. (3)

   The relationship between the stress arch structure and the position of the stopping line is evaluated as follows: ① ESUL: the stress concentration in the roof plate of the retracement channel of the underlying coal seam is the highest, because the overburden block of the extensive collapse zone acts directly on the roof plate of the retracement channel, resulting in relative difficulties in roof support. ② OUL: although the retracement channel roof pressure is minimal, the overlying rock structure has the potential for rotation or slippage instability. ③ ISUL-SD: the pressure on the roof of the retracement channel is small and the overburden structure is stable, which is conducive to the safe retraction of the support and not limited by the width of the end-mining coal pillar. ④ ISUL-LD: it is basically the same as the condition of stopping under the non-goaf, but is limited by the width of the end-mining coal pillar.

4. (4)

   The withdrawal technology of the self-digging retracement channel is adopted according to the different control requirements of the roof above the support and the roof of the retracement channel in the stopping area, through a comprehensive analysis of the pre-stress field simulation of the support scheme, a partition asymmetric support scheme is proposed. Field practice shows that the roof above the support is protected but not strengthened in the early partition strengthening support, and the strengthening effect of the combined anchor cable in the channel roof support is obvious. Therefore, this method realizes the safe and smooth withdrawal of the support.