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Article

Deformation Failure Characteristics and Maintenance Control Technologies of High-Stress Crossing-Seam Roadways: A Case Study

by
Zhengzheng Xie
1,
Zhe He
1,*,
Zhe Xiang
1,
Nong Zhang
1,2,
Jingbo Su
1,
Yongle Li
1 and
Chenghao Zhang
3
1
State Key Laboratory Coal Resource Mining and Safety Mining, Ministry of Education of China, School of Mines, China University of Mining and Technology, Xuzhou 221116, China
2
School of Civil Engineering, Xuzhou University of Technology, Xuzhou 221116, China
3
Laboratory of Geotechnics, Department of Civil Engineering, Ghent University, Zwijnaarde, 9052 Gent, Belgium
*
Author to whom correspondence should be addressed.
Appl. Sci. 2023, 13(7), 4442; https://doi.org/10.3390/app13074442
Submission received: 16 February 2023 / Revised: 28 March 2023 / Accepted: 28 March 2023 / Published: 31 March 2023
Browse Figures
Versions Notes

Abstract

:
The surrounding rock structure of the crossing-seam roadway is poor and is susceptible to anchorage failure phenomena, such as top plate sinking and convergence deformation under high ground stress. These issues can cause significant deformation of the surrounding rock over time, resulting in challenging engineering problems. To address this issue, we studied the failure modes and destabilization mechanisms of the surrounding rock in different crossing-seam roadways by field tests and numerical simulations. The results show that since the rock strata in these roadways are extremely unstable and highly susceptible to high horizontal stress, the weak surrounding rock presents the mode of full-section plastic failure. The roof is damaged more seriously than the floor and both walls. In this case, the basic anchorage layer in the original scheme is not thick and rigid enough to support these roadways. Thus, the surrounding rock deforms severely and persistently, which is one of the engineering failure characteristics. To solve this problem, a new scheme of “prompt thick-layer end anchorage + full-length lag grouting anchorage + secondary continuous reinforcement” was proposed based on the continuous roof control theory. According to the industrial test, this scheme can successfully control the long-term large deformation of the weak surrounding rock in crossing-seam roadways. Notably, the deformation of the top plate decreased by 56.65% and the deformation of the two walls decreased by 66.35%. Its design concept will provide important references for controlling the surrounding rock in similar soft rock roadways.

1. Introduction

Each year in China, tens of thousands of kilometers of new coal mine roadways are excavated [1]. The stable control of the surrounding rock in these roadways is directly related to the safe production and socioeconomic benefits of mines. Approximately 60% of discovered coal resources are located at depths of 800 m [2]. As fully mechanized mining methods continue to improve and mining intensity increases, shallow-buried resources are becoming increasingly depleted. As a result, deep coal mining has been increasingly adopted in many eastern mining areas. However, the failure mechanisms of deep soft rock roadways are different from those of shallow ones; and the deep rock mass also significantly differs from the shallow one in structure and mechanical property.
Compared to shallow-buried coals (rocks), deep-buried ones are affected by high seepage pressure, high in situ stress, large geothermal gradients, and intense mining disturbance in a complex coupled environment, in which the surrounding rock presents multiple nonlinear mechanical properties concerning soft rocks, such as high ground pressure, large-scale deformation, and difficulties supporting [3,4,5]. Additionally, there are a series of engineering responses in roadways, including obvious mine pressure, severe surrounding rock deformation, and significant rheological effects [6,7,8]. As the rock stress keeps rising, the rock mass becomes less stable under the influence of geologic structures. In this case, the surrounding rock in deep roadways tends to deform or even collapse under external disturbance. According to statistics, more than 30% of annually excavated roadways are deep soft rock ones that are restored at a rate of 70% [9,10], which seriously affects mining production and safety. Therefore, people have paid more attention to the stable control of these roadways [11,12].
The damage mechanism and control technology of soft rock roadways has been extensively studied, and in this process, a wealth of theoretical and practical experience is accumulated. In respect of failure mechanisms, He [13,14] systematically studied the characteristics of surrounding rock deformation in high-stress soft rock roadways based on elastic-plastic yield criterion, theoretical mechanics, engineering geology, and other classical theories and analyzed the cause of deformation, disaster evolution rules, and control techniques, which provided guidance for the support design of soft rocks. Mark [15] analyzed the failure mechanisms of the surrounding rock in deep soft rock roadways using UDEC software and proposed that destabilization was mainly caused by high horizontal stress. Tang [16] used RFPA software to simulate the process of surrounding rock failure in swelling rock tunnels under high-humidity conditions.
Wang [17] discussed the buckling deformation characteristics, damage evolution rules, and dynamic response mechanisms of the weak surrounding rock in deep tunnels under unloading disturbance by theoretical analysis and numerical simulation. Huang [18] analyzed the distribution and evolution characteristics of plastic damage zones in the surrounding rock of deep loop roadways with unequal pressure under high stress. Meng [19] researched the geotechnical characteristics of high-stress soft rock roadways using a viscoelastic rheological model and concluded that the surrounding rock featured the rheological properties of attenuated and stabilized deformation. Zhao [20] studied the local stress field of deep roadways and revealed their asymmetrical deformation failure mechanisms.
Yang et al. [21] explored the deformation failure characteristics of deep tunnels in compound strata. Zhao [22] researched the mechanical properties and spatial-temporal effects of deformation during the construction of deep soft rock tunnels by comparing the data obtained from field measurements and numerical simulations. Peng et al. [23] discussed the characteristics of unstable failure in kilometer-deep roadways under different pressure coefficients and analyzed the severe impact of high in situ stress on the stability of coupling-bearing strata. Wang [24] studied the failure scope, displacement trend, and stress distribution of elastic, plastic, and fracture zones in the surrounding rock before and after excavating deep soft rock roadways. Due to the complexity of geologic conditions and the diversity of engineering rock mass, there are various types of deep soft rock roadways whose failure mechanisms remain to be revealed.
The high-stress soft rock roadways are also studied at home and abroad concerning control techniques through engineering practice. The existing techniques are mainly divided into active reinforcement and passive load bearing, such as U-shaped steel supporting, pre-stressed bolt-cable supporting, rock grouting, and surrounding rock pressure releasing. Kang et al. [25] proposed the collaborative control concept of “three-actives” (active support with high-stress bolts—active grouting modification—active pressure relief by hydraulic fracturing) for high-stress and high-strength mining roadways in kilometer-deep mines. Ma et al. [26] discovered the butterfly-shaped plastic zone and put forward the theory of malignant expansion and the control theory of deep large-scale deformation. Wang et al. [27] developed the full-section deep-shallow coupled bolt grouting support for deep soft rock roadways.
Gao et al. [28] suggested supporting the high-stress soft rock roadways using concrete-filled steel tubes to avoid large deformation. Kang et al. [29] stabilized the soft rock roadways at a depth of −1180 m using “U-shaped steels + roof bolts and cables + floor grouting cables”. Feng and Zhang et al. [30] realized the gob-side entry retaining design in multiple mining areas by “anchoring + grouting + I-shaped beams + articulated steel bars + paste filling”. Jiang et al. [31] proposed an integrated support scheme for complementary control over seriously deformed high-stress roadways. Bai et al. [32] believed that the key to surrounding rock control in deep roadways was to strengthen the surrounding rock, transfer the high stress, and adopt reasonable support. Li et al. [33] established the new square steel confined concrete (SQCC) supporting system.
The surrounding rock in many deep soft rock roadways has been effectively controlled by existing support schemes, but these schemes, with their own geological characteristics, are not exactly applicable to other projects, especially in tectonic stress zones, fault fracture zones, and complex crossing-seam zones. As the primary type of roadway in coal mining operations, crossing-seam roadways typically cut through various rock formations at an incline to meet the requirements of the mining process. This results in highly complex conditions for the surrounding rock of the tunnel, particularly in deep areas where high stress is present. In these cases, it can be extremely challenging to maintain and control roadways that pass through weaker rock layers and their surrounding formations. The return air subinclined shaft in the Wuju Coal Mine is a typical deep soft rock roadway with high stress, which passes through dozens of strata composed of mudstone, coal, sandy mudstone, and sandstone. Additionally, this roadway is significantly affected by geologic structures, with a side-pressure coefficient of 1.4–1.5, complicating the occurrence conditions and stressful environments of its surrounding rock. However, it is only supported by the conventional “bolt-cable-mesh-spraying” method, which is so inadequate that the surrounding rock located in soft rock formations deforms within one month and cannot be stabilized through repeated restoration.
In this work, therefore, the return air subinclined shaft in the Wuju Coal Mine was selected as the object to systematically research the failure mechanisms and control techniques of deep crossing-seam roadways. By studying the failure modes and mechanisms of different crossing-seam roadways through theoretical analysis, numerical simulation, and field tests, we proposed the continuous roof control theory and technology for deep soft rock roadways in tectonic stress zones, based on which the new support scheme of “prompt thick-layer end anchorage + full-length lag grouting anchorage + secondary continuous reinforcement” was developed. In addition, a precise field test was conducted to evaluate the control effect of this scheme on the surrounding rock in crossing-seam soft rock roadways.

2. Engineering Overview and Issues

2.1. Geologic Setting

The tested roadway refers to the return air subinclined shaft in China’s Wuju Coal Mine. It is a development roadway extending from the first level (+600 m) to the second level (+300 m). In the process of excavation, this roadway passes through a total of 25 rock strata, including hard rock formations composed of fine-grained sandstone and coarse sandstone, and soft rock formations composed of coal, mudstone, and sandy mudstone. It is, therefore, defined as a typical crossing-seam roadway. This return air subinclined shaft with a dip angle of 20° is excavated parallel to the coal seam.
Table 1 lists the rock strata that the roadway mainly passes through during excavation. It can be seen that within a mileage of 108 m, the surrounding rock has high strength and good lithology; there is an interbedded formation of mudstone and sandstone between 108 m and 405 m, where the surrounding rock varies greatly in strength. The mudstone formation, located within 405~827.9 m, is a typical soft rock zone with low-strength surrounding rock. In this work, the tested roadway is a deep one at a burial depth of 799.1~918.9 m. The lithology distribution of the surrounding rock in the return air subinclined shaft is shown in Figure 1.
During excavation, the surrounding rock is primarily composed of mudstone that occurred in stratified and laminated forms, with joint fissures developed in this process, and the lithology is poor. The tested roadway passes through the 5-2 coal seam, where coals are brittle and inferior due to the low hardness of this seam. Moreover, under the obvious influence of geologic structures, it also passes through three tectonic fault zones. The in situ stress test indicates that horizontal stress greatly affects this roadway, showing a side-pressure coefficient of 1.4–1.5. In addition, there is an angle of about 25–43° between the maximum horizontal stress and the axis of this roadway.

2.2. Original Support Scheme

The return air subinclined shaft, which is 5.5 m wide and 4.95 m high, has a semicircular arch section of 23.98 m2 with vertical walls at a height of 2.2 m. The entire roadway is supported according to the same scheme, as shown in Figure 2.
① Primary support for the roof and both walls: The Φ20 mm × 2400 mm bolts were distributed at an inter-row spacing of 800 mm × 800 mm, with each row containing 16 bolts. The anchorage was about 1.3 m long.
② Reinforced support for the roof and both walls: One row of long cables was distributed in every two rows of bolts. The 1 × 7 stranded ordinary cables in the size of Φ17.8 mm × 7300 mm were arranged at an inter-row spacing of 1600 mm × 1600 mm, with each row containing seven cables. The anchorage was about 2.4 m long.
③ The arched anchor plates at a size of 150 mm × 150 mm × 10 mm and 300 mm × 300 mm × 16 mm were adopted for bolts and cables, respectively. The roof surface was further protected by a steel ladder beam and metal mesh. The metal mesh was welded with Φ6.5 mm steel bars.
④ After the support was completed, the exposed surface was sprayed with shotcrete (strength: C20) according to a thickness of 150 mm.

2.3. The Effect of Roadway Maintenance

The field investigation shows that the original support has a good maintenance and control effect on the roadway located in hard rock formations composed of sandstone or interbedded with mudstone and sandstone. In this scheme, the roadway only undergoes a small deformation. However, the surrounding rock deforms dramatically when it passes through the coal–mudstone interbedded formation. According to the field investigation in the return air subinclined shaft (Figure 3), the roadway deforms severely in local positions, which is manifested as roof subsidence and left wall heave, showing a long-term rheological feature. The surrounding rock is seriously fragmented at the heading face, developing many cracks. After the roadway is excavated, the monthly average deformation exceeds 400 mm on the roof to floor, and the contraction is over 300 mm on the wall to wall, which significantly affects the ventilation and transportation of the second level. Therefore, it needs a considerable amount of work to restore the surrounding rock.

2.4. Analysis of Geologic Occurrence and Maintenance Characteristics of the Roadway

(1) The surrounding rock shows complex stress conditions. The tested segment, which is buried at a depth of 799.1~918.9 m in the return air subinclined shaft, is a typical roadway with high in situ stress. It is significantly affected by folds when passing through three tectonic fault zones in excavation. The surrounding rock, with high tectonic stress, has a side-pressure coefficient of 1.4–1.5. Under these two kinds of stress, the surrounding rock is subject to asymmetrical deformation failure, manifesting as large deformation, long duration, and severe damage.
(2) The roadway passes through dozens of weak and unstable strata in the longitudinal direction. It is observed in the histogram of drilling data that the return air subinclined shaft longitudinally passes through a total of 25 rock strata, where the joint fissures of mudstone are developed and occur in an obvious stratified form. Additionally, since this is a multi-regional tectonic fault zone, the rock mass has low strength and poor stability.
(3) The original scheme is unreasonable. The anchorage is only 2.3 m thick in the primary support, which may easily result in overall deformation, and the cables (7.3 m) are too long to reinforce the anchorage layer in shallow strata. The effect of this collaborative reinforcement is so poor that a continuous roof structure cannot be formed. In addition, under unimproved pre-tightening force, the primary support weakens a wide range of shallow rock mass in active supporting. The basic anchorage layer is not rigid enough to maintain and control the surrounding rock under complex stress conditions for a long time. Therefore, there remains great space to optimize the support in the return air subinclined shaft.

3. Simulation Analysis on Failure Characteristics of the Crossing-Seam Roadway

3.1. Establishment of the Numerical Model

To study the deformation failure characteristics of the surrounding rock in different crossing-seam roadways, we established a numerical model (46 m × 45 m × 31.5 m) for the return air subinclined shaft using FLAC 3D software, as shown in Figure 4. The physical and mechanical parameters of each rock stratum are presented in Table 1. This model is a constitutive one and is in conformity with the Mohr–Coulomb yield criterion. The simulation model was meshed using the griddle plug-in in Rhino 6.0 software, and the mesh shape was selected as triangular and uniformly distributed. The mesh was automatically generated by the Rhino system, where the mesh width was set to 1.0 m. The model base plate and both sides are displacement boundary conditions. The model limits horizontal displacement on the side, vertical movement on the bottom, and stress boundaries on the upper surface. The top part of the model is the stress boundary. A vertical stress of 21.25 MPa was applied on the top (average burial depth: 850 m) to simulate a load of strata. With the side-pressure coefficient controlled at 1.5, the maximum horizontal stress that was set at 31.88 MPa formed an angle of about 30° with the axis of the roadway to simulate the tectonic stress of strata.
To reflect the failure characteristics of the surrounding rock in different crossing-seam roadways, we simulated the distribution laws of vertical stress, horizontal stress, and plastic zone in the following four cases (hereinafter referred to as Case 1, Case 2, Case 3, and Case 4). ① The roadway is completely located in hard rock formations, and the roof is composed of hard rocks; ② the roadway is completely located in hard rock formations, and the roof is composed of soft rocks; ③ the roadway is completely located in soft rock formations, and the floor is composed of hard rocks; and ④ the roadway is completely located in soft rock formations, and both the roof and floor are composed of soft rocks. The simulation results are shown in Figure 5.

3.2. Figures, Tables, and Schemes

3.2.1. The Distribution of Vertical Stress

Figure 6 illustrates the distribution of vertical stress in different crossing-seam roadways. When the roadway is completely located in hard rock formations (Case 1), the floor is a place where stress is mainly released, and the surrounding rocks on both sides show a phenomenon of stress concentration. However, as the roadway is excavated toward soft rock formations, this phenomenon is gradually attenuated. Meanwhile, stress is released in a wider range on the floor, resulting in the release on the roof, which implies that a large amount of stress is released from soft rock formations. Then, when the roadway is completely located in these formations, stress is released to the maximum extent. At this point, the largest stress-releasing areas are formed on the roof and floor, which indicates severer deformation in both positions.

3.2.2. The Distribution of Horizontal Stress

The distribution of horizontal stress in different crossing-seam roadways is shown in Figure 7. When the roadway is completely located in hard rock formations, stress is concentrated on both sides and released on the floor and right wall. As the roadway is excavated into soft rock formations (Cases 2 and 3), the phenomenon of stress concentration disappears from both sides but appears on the roof, causing obvious tensile failure. However, the floor remains the place where stress is mainly released. Ultimately, when the roadway fully enters soft rock formations, the floor becomes the area with the largest stress release, while the roof and right wall experience stress release to a lesser extent.

3.2.3. The Distribution of the Plastic Zone

Figure 8 shows how the plastic zone is distributed in different crossing-seam roadways. When the roadway is located in hard rock formations, the plastic zone occupies a small area and primarily distributes on the roof. As the roadway is excavated toward soft rock formations (Cases 2 and 3), the plastic zone gradually expands, which mainly occurs on the roof and rarely occurs on the floor and both walls. When the roadway fully enters soft rock formations, the plastic zone is obviously expanded on the roof, floor, and walls. Under such circumstances, the roadway presents a state of full-section plastic development, in which the surrounding rock is extremely unstable.

3.2.4. Deformation on the Roadway Surface

The amount of deformation in the surrounding rock of different crossing-seam roadways is presented in Figure 9. In the above four cases, the roof-to-floor deforms more seriously than the wall-to-wall, indicating that the horizontal stress mainly works on the former. Further, based on the previous elaboration in the plastic zone, it is deduced that this stress primarily affects the deformation failure on the roof. The amount of deformation has little difference between Case 1 and Case 2, which implies that the lithology of the roof has a small impact on surrounding rock deformation when the roadway is located in hard rock formations. Nevertheless, the amount is significantly increased in Case 3 and reaches its maximum value in Case 4, suggesting that horizontal stress has a clear impact on soft rocks.
To sum up, when the roadway is completely located in hard rock formations, or when it is located in these formations while the roof is in soft ones, stress is concentrated on both sides, and the surrounding rock slightly deforms. When the roadway is located in soft rock formations while the floor is in hard ones, the plastic zone is significantly expanded on the roof, which exacerbates the deformation failure of the surrounding rock. When the roadway fully enters soft rock formations, this zone becomes much larger than that in Cases 1~3. At this point, the roadway presents a state of full-section plastic development, and the roof has a much larger plastic zone than the floor and both walls. This indicates that horizontal stress can cause greater damage to the roof than to the floor and both walls, which is increasingly obvious under reduced surrounding rock strength. The roadway completely located in soft rock formations is damaged more severely than that in hard ones. In this case, the support design shall be emphasized. That is, the basic anchorage layer shall be thickened and lengthened accordingly to exceed the critical level, thus stabilizing the roadway persistently.

4. Continuous Roof Control Theory and Technology for High-Stress Soft Rock Roadways

After the high-stress roadway is excavated, the rock mass undergoes flexural deformation toward the free face in the stress rebalancing process. The varying forced states and evolution courses can cause different degrees of failure, flexural deformation, and horizontal displacement to the peripheral rock mass, which is known as the discontinuous deformation failure that is intensified under reduced surrounding rock strength. The evolution of this failure is illustrated in Figure 10.
In the process from the roadway excavation to stabilization, the peripheral rock mass deforms seriously in shallow strata, where the deformation gradually attenuates and finally terminates as the depth increases. The fissure development in the rock mass surrounding the roadway generally corresponds to the relative deformation of rock layers. This is typically characterized by the development of shallow fissures and the gradual disappearance of deep rock mass fissures. Hence, if the basic anchorage layer constructed by bolts has an unreasonable thickness, the long-term deformation of the surrounding rock cannot be effectively controlled even if high pretightening force is applied to the supporting structure, which is manifested as the evolution from surface fissures to deep cracks. Then, as the surrounding rock is progressively fractured in a wider range under this support, these cracks can easily extend outside the anchorage layer constructed by short bolts, which obviously reduces the rock mass stability on the roof. In this instance, the surrounding rock deforms severely under the condition that bolts are not broken but have reduced bearing capacity. On the contrary, with the continuously deepened high pre-stressed support anchoring into the surrounding rock mass, the sensitivity of the anchoring rod to the relative displacement of the mass is constantly strengthened. When the basic support structure penetrates the transverse fissure zone, the anchoring structure can effectively mobilize a larger range of rock mass to construct a load-bearing structure, thus achieving long-term stable control and improving the ability to resist dynamic load disturbance.
In addition to the anchoring thickness formed by the support structure, the magnitude of the pre-stressed force of the anchoring rod plays a decisive role in the stability of the anchoring rock beam of the top plate. When the force is large enough, the displacement of the top plate within the length range of the anchoring rod, especially within the effective range of the pre-stressed force, is controlled. Therefore, the key point of continuous strengthening control is to construct a thick anchoring structure using a long anchoring rod and apply a relatively large pre-stressed force to fully utilize the coordinated deformation between rock layers. A large number of field measurements have shown that the formation of the top plate eliminates the load increase in the anchoring rod or cable in the later stage, which means that the rod itself does not need to have any other function in addition to maintaining the pre-stressed force.
Therefore, the anchored rock beams that are thick enough are constructed with pre-stressed bolts to eliminate the abscission layer and two types of destabilization, thus realizing the bidirectional linkage within the same layer and among different layers. A thick anchorage-bearing layer is built using flexible supporting rods [34,35], as shown in Figure 11. This layer spans the shallow damage zone to provide a larger load-bearing circle for the surrounding rock so that the deformation in shallow strata can be restrained through the long-term stabilization of the less damaged surrounding rock in deep strata. In addition to this thick-layer anchorage, a continuous roof structure is synergistically constructed with secondary reinforcement cables based on the surrounding rock conditions. Thus, the multi-layer anchored rock beams can bear various superimposed loads to continuously transfer stress between these beams, thus avoiding the generation and propagation of cracks in the abscission layer. In this way, the supporting structure becomes more resistant to multiple stress perturbations.
On the basis of the thick-layer anchorage with high prestress, lag grouting is performed on hollow cables to integrate them with the surrounding rock. The initial fissures within the anchorage zone are cemented and plugged to increase the sensitivity of cable rods to surrounding rock deformation, thus further inhibiting crack propagation and deformation. Meanwhile, the basic thick-layer anchorage structure is improved in rigidity to strengthen the constraint on the shallow surrounding rock. Since this structure has high strength and rigidity, it can address the lack of pretightening force in traditional full-length anchorage supports and the insufficient control over fractured rock mass in traditional end anchorage supports.

5. Industrial Field Test

5.1. New Support Scheme

Based on the above analysis and relevant case studies, the two-level reinforcement was designed to achieve a high-quality continuous and transboundary multilevel anchorage for the soft rock roadway in the return air subinclined shaft. The high quality (high integrity, high strength, and high rigidity) in the primary basic support is realized by reasonable anchorage thickness + prompt end anchorage support + strengthened constraint on shallow strata + full stress release + full-length lag grouting anchorage, which ensures low damage to the surrounding rock in the shallow load-bearing circle.
Primary support is continuously reinforced by secondary support through synergistic anchorage thickness + lengthened strong anchorage to stabilize the roadway for a long time. The basic and reinforced load-bearing circles in primary and secondary supports are 3.5 m thick (rod length: 3.8 m) and 5.5 m thick (rod length: 5.8 m), respectively, which guarantee the continuity of stress within the two circles. The new support scheme was proposed based on the treatment concept of “prompt thick-layer end anchorage + efficient full-length lag anchorage + secondary continuous reinforcement”, as shown in Figure 12a.
(1) Support scheme.
① Primary basic support for the roof: The high-strength grouting cables (Φ21.6 mm) at a length of 3.8 m (Figure 12b) were arranged on the roof at an inter-row spacing of 800 mm × 800 mm, with each row containing 11 cables. The pretightening force was not lower than 150 kN, and a steel ladder beam was constructed for reinforcement. To enhance the shearing resistance at humeral angles and the protective effect on the roadway surface, 500 right-handed fine-tooth threaded steel bolts (Φ22 mm) at a length of 2.4 m (Figure 12c) were distributed at an inter-row spacing of 1200 mm × 1600 mm, with each row containing five bolts.
② Secondary reinforced support for the roof: The roof was further supported by Φ21.8 mm × 5800 mm ordinary cables at a length of 5.8 m, which were arranged at an inter-row spacing of 1200 mm × 1600 mm, with each row containing seven cables. The pretightening force was not lower than 180 kN.
③ Wall support: Both walls were supported by four right-handed fine-tooth threaded steel bolts (Φ22 mm × 2400 mm) and two high-strength grouting cables (Φ21.6 mm × 3800 mm) at an inter-row spacing of 800 mm × 800 mm. Bolts were distributed vertically to both walls, with the pretightening torque not lower than 300 N·m; cables were arranged at a depression angle of 15° near the floor, with the pretightening force not lower than 150 kN.
④ As diagrammed in Figure 12d, anchor plates at a size of 150 mm × 150 mm × 10 mm were used for bolts; the large arched ones with high strength in a size of 300 mm × 300 mm × 20 mm were adopted for ordinary cables and grouting ones used on the roof. Steel-bar mesh (Φ6.5 mm × 2100 mm × 900 mm) and rhombic metal mesh (Φ10 mm × 2400 mm × 1000 mm) were used to reinforce the roof and both walls, respectively.
⑤ Upon the completion of support, the exposed surrounding rock was sprayed in a thickness of about 30~50 mm. The grouting cables were grouted with cement-based non-shrinkage grouting materials for the full length at a lag distance of 30~50 m under a grouting pressure of 2–3 MPa. After this operation, the surrounding rock was resprayed with shotcrete (strength: C20) at a thickness of 150 mm.

5.2. Field Observation and Maintenance Effect

The mine pressure at the tested segment of the return air subinclined shaft under the new support was monitored by the cross-shaped point distribution method using the borehole imager and bolt-cable dynamometer. The results were compared with those in the original scheme as follows.
(1) Borehole observation of the surrounding rock.
The borehole images observed on the roof under the original support are presented in Figure 13. It can be seen that annular cracks are found at 0.47 m, 0.72 m, 0.91 m, 1.27 m, 1.48 m, 1.59 m, 2.14 m, 2.16 m, 2.75 m, 2.84 m, and 5.01 m; vertical cracks are developed at 1.0 m, 1.21 m, 1.91 m, 2.0 m, 2.31 m, and 4.85 m, which can easily result in the formation of abscission layers or even fracture zones. Abscission layers are formed at 2.38 m, 2.59 m, 2.65 m, and 2.84 m; longitudinal holes are observed within 1.22~1.34 m. The continuous fracture zones are distributed from 2.38 m to 2.43 m, with the longest one reaching 5 cm, where the cracks are developed at a maximum depth of 5.01 m. In the original scheme, the surrounding rock on the roof has poor integrity, developing numerous cracks, abscission layers, and continuous fracture zones. The cracks, with a maximum depth of 5.01 m, tend to extend deeper. This scheme has an extremely poor maintenance and control effect on the roadway.
According to the borehole images observed on the roof under the new support (Figure 14), minor and annular cracks are formed at 0.13 m, 0.68 m, 0.87 m, and 1.19 m, and there are cracks filled with slurry at 1.9 m. This indicates that the basic thick-layer anchorage can inhibit the generation and propagation of initial rock cracks and guarantee the strength of the initial support. Meanwhile, cracks are plugged with slurry to rebond the surrounding rock into an integral whole, which not only improves the overall strength of the fractured rock mass but also significantly enhances roadway stability.
(2) Axial force monitoring of bolts and cables.
Figure 15 shows how the axial force of bolts and cables evolves in the original scheme and how the axial force of bolts, cables, and grouting cables evolves in the new scheme. According to the observation on mine pressure, the axial force of bolts and cables includes two parts: one refers to the pre-tightening force applied during the installation of bolts and the other is the axial tension on bolt rods caused by surrounding rock deformation. Therefore, how the rock mass evolves in deep and shallow strata can be obtained by observing the change of this axial force.
After the pre-tightening operation is finished, the axial force of bolts in the original scheme is 48 kN. Then, the value surges to 178.22 kN and is finally stabilized at 177.25 kN, as shown in Figure 15a. The axial force of cables is 105.96 kN after this operation, and then it gradually rises over time, during which there is a phenomenon of sudden increase and decrease caused by the fracture of rock mass on the surface. It can be observed that the axial force of bolts rises, which indicates the severe deformation of shallow rock mass, manifesting as the formation of abscission layers and the propagation of cracks. By contrast, the axial force of cables increases gradually, which implies that the deep rock mass expands and deforms slightly at a low speed. These results are consistent with the image data obtained from borehole observation.
Figure 15b shows that the axial force of bolts in the new scheme is 68.59 kN after pre-tightening. Then, it gradually rises to 103.03 kN. The axial forces of ordinary and grouting cables are 163.09 kN and 106.16 kN, respectively, after pre-tightening. Then, the values slowly increase to 185.55 kN and 174.79 kN, respectively. The slow growth of this force in bolts, cables, and grouting cables proves the gradual expansion and small deformation of deep and shallow surrounding rocks under the new support. Additionally, there is no large-scale deformation failure on the roof and both walls, which indicates that the cracks are markedly inhibited from propagating in the surrounding rock, only causing a small amount of deformation. The roadway is effectively controlled and maintained by the new support.
(3) Deformation monitoring on the roadway surface.
The deformation characteristics under original and new supports are compared in Figure 16. The roof-to-floor deforms are 438.5 mm and 190.1 mm in the original and new schemes, respectively, and the wall-to-wall contracts are 384 mm and 129.2 mm, respectively. In both schemes, the roof-to-floor deforms more seriously than the wall-to-wall, which is in line with the results of the numerical simulation and the evolution of the axial force of bolts. The total deformation of these two types in the original scheme is much higher than in the new one. The photo taken on-site also verifies that the roadway deforms slightly and can be effectively controlled under the new support. This indicates that the new scheme of “prompt thick-layer end anchorage + efficient full-length lag anchorage + secondary continuous reinforcement” can realize the safe maintenance and control of high-stress crossing-seam soft rock roadways.

6. Conclusions

(1) The return air sub-inclined shaft in the Wuju Coal Mine is a typical high-stress roadway that crosses several weak, soft rock formations. Under the combined influence of in situ and horizontal stress, the surrounding rock located in soft rock formations undergoes long-term and large-scale asymmetrical deformation failure. The roof-to-floor deforms at 438.5 mm in the original schemes, and the wall-to-wall contracts are 384 mm.
(2) The numerical simulation results show that when the roadway is completely located in soft rock formations, the plastic zone is much larger than that in the first three cases, and the roadway presents a state of full-section plastic development. The surrounding rock on the roof is damaged more seriously than on the floor and both walls. Given this, the basic anchorage layer shall be thickened to reinforce the support.
(3) For the crossing-seam roadway, the treatment concept of “prompt thick-layer end anchorage + full-length lag grouting anchorage + secondary continuous reinforcement” was proposed based on the continuous roof control theory. According to the monitoring data about mine pressure, the new support enhanced the surrounding rock’s stability, the deformation of the top plate decreased by 56.65%, and the deformation of the two walls decreased by 66.35%. The roadway presents the mode of gradual expansion and small deformation in both shallow and deep strata, which proves that the new scheme can successfully control the long-term large deformation of this roadway.

Author Contributions

Methodology, Z.X. (Zhe Xiang) and Z.X. (Zhengzheng Xie); formal analysis, Z.H. and Z.X. (Zhengzheng Xie); software, Z.H. and J.S.; investigation, Y.L., C.Z., N.Z. and Z.X. (Zhengzheng Xie); writing—original draft, Z.X. (Zhengzheng Xie); writing—review and editing, Z.X. (Zhengzheng Xie), Z.X. (Zhe Xiang) and Z.H. All authors have read and agreed to the published version of the manuscript.

Funding

This study was funded by the National Natural Science Foundation of China (grant No. 52104104).

Institutional Review Board Statement

Not applicable.

Informed Consent Statement

Not applicable.

Data Availability Statement

If relevant data are required, please contact the corresponding author by email.

Conflicts of Interest

The authors declare no conflict of interest.

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Figure 1. Lithology distribution of the surrounding rock in the return air subinclined shaft.
Figure 1. Lithology distribution of the surrounding rock in the return air subinclined shaft.
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Figure 2. Cross-sectional diagram for the bolt-cable coordinated support in the return air subinclined shaft.
Figure 2. Cross-sectional diagram for the bolt-cable coordinated support in the return air subinclined shaft.
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Figure 3. The effect of roadway maintenance under original support: (a) location I; (b) location II.
Figure 3. The effect of roadway maintenance under original support: (a) location I; (b) location II.
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Figure 4. Numerical model.
Figure 4. Numerical model.
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Figure 5. Four rock strata: (a) Case 1; (b) Case 2; (c) Case 3; (d) Case 4.
Figure 5. Four rock strata: (a) Case 1; (b) Case 2; (c) Case 3; (d) Case 4.
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Figure 6. The distribution of vertical stress in the roadway: (a) Case 1; (b) Case 2; (c) Case 3; (d) Case 4.
Figure 6. The distribution of vertical stress in the roadway: (a) Case 1; (b) Case 2; (c) Case 3; (d) Case 4.
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Figure 7. The distribution of horizontal stress in the roadway: (a) Case 1; (b) Case 2; (c) Case 3; (d) Case 4.
Figure 7. The distribution of horizontal stress in the roadway: (a) Case 1; (b) Case 2; (c) Case 3; (d) Case 4.
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Figure 8. The distribution of the plastic zone in the roadway: (a) Case 1; (b) Case 2; (c) Case 3; (d) Case 4.
Figure 8. The distribution of the plastic zone in the roadway: (a) Case 1; (b) Case 2; (c) Case 3; (d) Case 4.
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Figure 9. Deformation law on the roadway surface.
Figure 9. Deformation law on the roadway surface.
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Figure 10. The evolution of discontinuous deformation failure in the surrounding rock.
Figure 10. The evolution of discontinuous deformation failure in the surrounding rock.
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Figure 11. Schematic diagram for the continuous roof control.
Figure 11. Schematic diagram for the continuous roof control.
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Figure 12. Support scheme and supporting components. (a) New support scheme; (b) The high-strength grouting cables; (c) Bolt; (d) Anchor plates.
Figure 12. Support scheme and supporting components. (a) New support scheme; (b) The high-strength grouting cables; (c) Bolt; (d) Anchor plates.
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Figure 13. Borehole observation under the original support. (a) 0.47 m; (b) 0.72 m; (c) 0.91 m; (d) 1.27 m; (e) 1.48 m; (f) 1.59 m; (g) 1.91 m; (h) 2.16 m. (i) 2.59 m; (j) 2.65 m; (k) 2.75 m; (l) 3.45 m. (m) 3.92 m; (n) 5.01 m.
Figure 13. Borehole observation under the original support. (a) 0.47 m; (b) 0.72 m; (c) 0.91 m; (d) 1.27 m; (e) 1.48 m; (f) 1.59 m; (g) 1.91 m; (h) 2.16 m. (i) 2.59 m; (j) 2.65 m; (k) 2.75 m; (l) 3.45 m. (m) 3.92 m; (n) 5.01 m.
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Figure 14. Borehole observation under the new support. (a) 0.13 m; (b) 0.68 m; (c) 0.87 m; (d) 1.19 m; (e) 1.90 m.
Figure 14. Borehole observation under the new support. (a) 0.13 m; (b) 0.68 m; (c) 0.87 m; (d) 1.19 m; (e) 1.90 m.
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Figure 15. Axial force monitoring of bolts and cables. (a) Original scheme; (b) new scheme.
Figure 15. Axial force monitoring of bolts and cables. (a) Original scheme; (b) new scheme.
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Figure 16. Deformation characteristics on the roadway surface. (a) Amount of the surface deformation; (b) image of the surface deformation.
Figure 16. Deformation characteristics on the roadway surface. (a) Amount of the surface deformation; (b) image of the surface deformation.
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Table
Table 1. Lithology distribution of the surrounding rock in the return air subinclined shaft within different mileages.
Table 1. Lithology distribution of the surrounding rock in the return air subinclined shaft within different mileages.
Rock TypeElevation/mBurial Depth/mMileage/mLithology DescriptionRemarks
Fine-grained sandstone+585.6653.552.60The main component is quartz, and the rock mass has high hardness.Sandstone formation; total mileage: 108.12 m
Coarse sandstone+566.3672.8108.12The main component is quartz, and the rock mass has high hardness.
Mudstone+553.8685.3145.13The stratum, interbedded with slack coals, is relatively broken, and the rock mass has low integrity.Mudstone–sandstone interbedded formation; total mileage: 296.88 m
Fine-grained sandstone+540.0699.1184.75The main component is quartz, and the rock mass has high hardness.
Mudstone+520.0719.1241.95The stratum is developed in a stratified form, and the rock mass has low integrity.
Sandy mudstone+507.2731.9289.70The stratum is developed in a stratified form, and the rock mass is hard with moderate integrity.
Fine-grained sandstone+494.6744.5318.58The main component is quartz, and the rock mass has high hardness.
Mudstone+479.0760.1364.28The stratum is subject to water softening, and the rock mass has low integrity.
Fine-grained sandstone+464.8774.3405.78The main component is quartz, and the rock mass has high hardness.
Mudstone+440.0799.1478.30The stratum is developed in a stratified form, and the rock mass has low integrity.Coal–mudstone interbedded formation; total mileage: 422.9 m
Sandy mudstone+422.5816.6529.50The rock mass is hard, with moderate integrity.
5-2 coal seam+420.0819.1538.70The coals are inferior and brittle.
Carbonaceous mudstone+409.2829.9568.50The stratum, developed in an oblique bedding form, is relatively broken.
Sandy mudstone+380.8858.3651.30The rock mass is hard, with moderate integrity.
Mudstone+320.2918.9828.60The stratum, with clear stratification, is relatively broken.
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Xie, Z.; He, Z.; Xiang, Z.; Zhang, N.; Su, J.; Li, Y.; Zhang, C. Deformation Failure Characteristics and Maintenance Control Technologies of High-Stress Crossing-Seam Roadways: A Case Study. Appl. Sci. 2023, 13, 4442. https://doi.org/10.3390/app13074442

AMA Style

Xie Z, He Z, Xiang Z, Zhang N, Su J, Li Y, Zhang C. Deformation Failure Characteristics and Maintenance Control Technologies of High-Stress Crossing-Seam Roadways: A Case Study. Applied Sciences. 2023; 13(7):4442. https://doi.org/10.3390/app13074442

Chicago/Turabian Style

Xie, Zhengzheng, Zhe He, Zhe Xiang, Nong Zhang, Jingbo Su, Yongle Li, and Chenghao Zhang. 2023. "Deformation Failure Characteristics and Maintenance Control Technologies of High-Stress Crossing-Seam Roadways: A Case Study" Applied Sciences 13, no. 7: 4442. https://doi.org/10.3390/app13074442

APA Style

Xie, Z., He, Z., Xiang, Z., Zhang, N., Su, J., Li, Y., & Zhang, C. (2023). Deformation Failure Characteristics and Maintenance Control Technologies of High-Stress Crossing-Seam Roadways: A Case Study. Applied Sciences, 13(7), 4442. https://doi.org/10.3390/app13074442

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