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Article

Numerical Simulation of Deformation and Failure Mechanism of Main Inclined Shaft in Yuxi Coal Mine, China

by
Fan Wu
1,
Yueping Qin
1,2,
Hao Xu
1,3,*,
Fengjie Zhang
1 and
Xiangyu Chu
1
1
School of Emergency Management and Safety Engineering, China University of Mining & Technology, Beijing 100083, China
2
Beijing Key Laboratory for Precise Mining of Intergrown Energy and Resources, China University of Mining & Technology, Beijing 100083, China
3
Petroleum Systems Engineering, Faculty of Engineering and Applied Science, University of Regina, Regina, SK S4S 0A2, Canada
*
Author to whom correspondence should be addressed.
Appl. Sci. 2022, 12(11), 5531; https://doi.org/10.3390/app12115531
Submission received: 6 May 2022 / Revised: 24 May 2022 / Accepted: 26 May 2022 / Published: 30 May 2022
(This article belongs to the Special Issue Safe and Efficient Exploitation and Utilization of Deep Earth Resources)
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Abstract

:
Disturbance stresses can cause deformation and damage to a tunnel’s rock, potentially threatening the mine’s safety. This paper investigates the effects of disturbance damage on the main inclined shaft due to the excavation of an electromechanical chamber in a deep inclined shaft at Yuxi Mine. Specifically, a numerical model was constructed using Midas GTX NX and Fast Lagrangian Analysis of Continua in Three Dimensions (FLAC3D) to match the actual engineering conditions, and to reveal the stresses and deformations in the surrounding rock of the main inclined shaft before and after the excavation of the main inclined shaft, the electromechanical chamber and the head chamber. The results revealed that the surrounding rock stress around the main inclined shaft is significantly influenced by excavation disturbance. The bottom bulge occurred due to the unstable vertical and shear stresses in the bottom coal bed moving into free space. After the excavation of the electromechanical chamber, the maximum displacement of the floor can be increased from 0.35468 m to 0.64301 m, nearly doubled, and a large area of surrounding rock deformation occurs in the inclined shaft falling roadway. Affected by excavation disturbance, the maximum deformation of floor can reach 1.06 m, with a wide fluctuation range. The main area of damage to the surrounding rock was identified, except for the main inclined shaft, which occurred near the intersection of the inclined shaft and the drop level location. This area is mainly affected by superimposed tensile stress damage, prone to large area floor heave and spalling. The research content is expected to provide certain theoretical support in taking measures to deal with the deformation and failure of the surrounding rock in a main inclined shaft.

1. Introduction

With the increasing depletion of shallow mineral resources, people are aiming at the development of deeper mineral resources [1,2,3,4]. The mining depth of coal in China shows an increasing trend each year. At present, the average mining depth has reached 700 m [5]. However, deeper mining faces more complex engineering geological conditions. In the process of coal mining, the deep surrounding rock is affected by the disturbance of high ground stress, and the frequency and intensity of disasters are increasing [6,7,8]. Especially in the deep excavation of the inclined shaft and chamber, it is easy for the surrounding rock to cause stress superposition or redistribution, which will lead to excessive floor heave and two-side displacement, resulting in large deformation and instability of the roadway. Accidents caused by large deformation and instability of the roadway account for a large proportion of the total accident coal, resulting in serious casualties and economic losses, which seriously restricts the safe and efficient mining of coal resources [9]. Studying the deformation and failure characteristics and mechanism of surrounding rock in a deep roadway is the basis for solving this kind of problem, which has important theoretical guiding significance and engineering application value to ensure the safe and efficient mining of a deep mine.
At present, there are many studies on the damage and failure characteristics of surrounding rock in deep engineering. According to the existing research results, in-situ stress is the basic force causing the deformation and failure of surrounding rock, and it plays an important role in controlling the deformation and failure of the surrounding rock [10,11]. Some scholars have used theoretical calculation to study the deformation characteristics of coal seams and roadways under water pressure [12,13,14]. However, most of these studies are to analyze the shaft accidents under non-mining conditions and lack investigation of the deformation characteristics of surrounding rock in the normal excavation process of the roadway or chamber. In underground engineering such as a coal mine roadway, due to the unloading disturbance after roadway or chamber excavation, the original stress balance state of a rock mass formed by long-term geological movement is destroyed, resulting in the redistribution of surrounding rock stress [9,15]. In the process of constant stress adjustment, the surrounding rock of the roadway will slide along the joint fissures and open holes, resulting in large deformation and instability of the surrounding rock [16]. Moreover, with the increase of mining depth, the large deformation of the roadway’s surrounding rock caused by excavation and unloading becomes more and more serious. Because the surrounding rock environment of the excavated roadway is complex, the theoretical calculation is difficult, and there are many influencing factors on engineering problems, it is a better solution to establish a physical model for numerical simulation. The numerical simulation method reveals the micro and macro mechanical behavior of the large deformation instability process caused by eccentric stress after roadway excavation [17,18,19,20,21]. For example, Li et al. conducted numerical simulation of roadways with different in-situ stress and rock mass structure based on a two-dimensional particle flow program (PFC2D) [22]. Zhang et al. established the equivalent post-peak strain softening model of jointed rock mass, and carried out numerical simulation by using FLAC3D software [4]. Hou and Yang created a physical model of the underground roadway in the horizontal stratum and monitored the surface displacement of the model using 2D digital image correlation technology [23]. Song et al. established a three-dimensional geological model of the mining area and numerically analyzed the influence of the combined mining of No. 3 and No. 9 coal in a steep coal seam on the stress and deformation of the surrounding rock of the main inclined shaft [11]. Wang et al. used the discrete element numerical model to reveal the asymmetric large deformation law, stress distribution characteristics and failure mechanism of jointed surrounding rock under excavation unloading [9]. Chen et al. divided the initial heterogeneous excavation damage zone of surrounding rock according to the mining and operation process, and studied the stability of CAES cavern in an abandoned mine tunnel [24]. In order to comprehensively understand the stability of a deep cavern after excavation disturbance, Zha et al. systematically proposed an indoor simulation method of a surrounding rock test under excavation disturbance [25]. In fact, coal seam mining will inevitably affect the mining of nearby coal seams, while the overburden undergoes secondary disturbance-induced fracture, the stress distribution is quite complex and the failure form of roadway-surrounding rock is also very complex [15,26]. Unfortunately, even though scholars have carried out a lot of theoretical analysis, physical models and field investigation, the disasters caused by roadway instability are still very common in deep mining engineering [22,27]. The deformation mechanism of roadway-surrounding rock under a deep dynamic load has not been fully studied, and the disaster mechanism is not clear. The factors affecting the rock stability around the roadway are complex. It is necessary to further explore the stress distribution characteristics and failure forms of the surrounding rock after excavating the roadway or chamber according to the specific geological conditions of the coal mine.
In view of the above scientific problems in roadway excavation, taking the geological conditions of the main inclined shaft of the Yuxi coal mine as the background, this paper uses Midas three-dimensional modeling software and the FLAC3D finite element analysis program to establish a numerical model matching with the surrounding rock of the actual roadway engineering, to simulate the stress disturbance and strain displacement of the surrounding rock of the main inclined shaft under the action of excavation unloading, and to reveal the asymmetric deformation law, stress distribution characteristics and failure mechanism of the surrounding rock. The main purpose is to provide a reliable scientific theoretical basis for the reasonable and stable control of the surrounding rock of a deep roadway and to effectively ensure the safe production of the coal mine.

2. Engineering Overview and Research Background

2.1. Geographical Location

The Yuxi coal mine is located in the southeast of Fanzhuang census area in Jincheng City, Shanxi Province, China, and its geographical location is shown in Figure 1. The field area is 26.1572 km2, the reserves of No. 3 coal resources is 26.7 million tons and the service life is 50.7 years, that is, the service life of the main inclined shaft in Yuxi Coal Mine is nearly 51 years. The normal use of the main inclined shaft directly affects the development and utilization of resources in the region.
The main inclined shaft of Yuxi Coal Mine started construction in 2012 and was installed and completed in 2013. The shaft is constructed by blasting excavation technology and supported by bolt + anchor cable + shotcrete. Due to the deep mining depth of the mine, the roof and floor strata of the coal seam are soft. After the completion of the wellbore installation, affected by the construction of the mechanical and electrical chamber at the bottom of the well, the surrounding rock of the tail drum section and the falling section has different degrees of deformation and failure. Tiandi Science and Technology Co., Ltd., has carried out the governance of this section of the roadway. The governance mainly adopts the shallow deep hole grouting and anchor cable support scheme for the two sides and the roof, and the borehole grouting reinforcement for the floor. Subsequently, affected by the tunneling of the head chamber of the belt roadway, the roadway has experienced different degrees of deformation and failure within the range of about 160 m from the bottom of the well. Among them, the deformation and failure of roadway in the range of about 100 m from the pipe mouth to the bottom of the well are relatively serious. The recent field investigation found that the roadway was still deformed and damaged, and the degree of deformation and damage had affected the safety and normal use of the roadway. The position relationship of the three roadways is shown in Figure 2.

2.2. Engineering Geological Characteristics

The elevation of the horizontal section of the main inclined shaft in Yuxi Coal Mine is +320, and the deepest distance is 582 m from the surface. Therefore, the original stress of the wellbore failure section itself is larger. The lithology of the main inclined shaft is mainly composed of mudstone and sandy mudstone, with a small amount of medium–fine sandstone, siltstone and limestone. The local surrounding rock shows obvious mechanical properties of soft rock. The main aquifers in the mine field mainly include karst fissure aquifers in the Middle Ordovician system, Taiyuan Formation aquifers in the upper Carboniferous system, Shanxi Formation aquifers in the lower Permian system, sandstone fissure aquifers in the upper Permian system, bedrock weathering zone aquifers, and Quaternary loose sand and gravel aquifers.
The space of the direct water-filled aquifer in the roof of the No. 3 coal seam is dominated by sandstone fractures, which belong to the aquifer of the lower Permian Shanxi Formation. This layer is a clastic rock fractured aquifer, and there is no surface spring outflow in the mine field. The aquifer is mainly composed of medium–fine sandstone: thickness 1.51~15.02 m, average 8.97 m. In the drilling process, the maximum consumption is 0.242 m3/h, the minimum is 0.026 m3/h, generally, 0.161 m3/h. The water level has no obvious change, the unit water inflow is 0.001~0.00281 L/s·m, the permeability coefficient is 0.0073~0.0105 m/d, which is a weak water-rich aquifer. The water quality is Cl-K + Na or Cl-HCO3-K + Na. The safety of mining under pressure in the No. 3 coal seam is evaluated. The water inrush coefficient of Ordovician limestone karst water in the pressure area is 0.025~0.047 mpa/m. Normal sections will not cause Ordovician limestone water inrush, and there is water inrush in structural water diversion sections. Water prevention (isolation) measures shall be taken during excavation in the structural water diversion section.

2.3. Research Background

Due to the mining disturbance or the influence of mining, the stress state of the surrounding rock changes and the mechanical properties of the surrounding rock changes in the support process, resulting in the deformation of the roof and floor of the roadway and the two sides of the rock mass. moving to the free space (in the roadway) and the floor heaving upward, which is called the floor heave. In general, a small amount of floor heave does not endanger the stability of the roadway, because the bottom drum in this range of magnitude has little effect on underground transportation and ventilation. Usually, when the floor heave of the roadway is less than 200 mm, special prevention measures are not needed. However, because the main inclined shaft of Yuxi Coal Mine has the dual role of transporting coal and pedestrians, a large amount of floor heave will affect its normal use, thereby hindering production; and because the floor is the foundation of the roadway, severe floor heave will lead to the instability of the whole roadway and a disaster.
The main inclined shaft of Yuxi Coal Mine is an important operation channel of the life main line of the whole mining area, which is directly related to the safe and efficient production of coal resources in the whole mining area. The main inclined shaft adopts a blasting excavation technology construction and the ‘anchor + anchor + shotcrete’ combined support method. At present, due to the influence of the construction disturbance of the mechanical and electrical chamber at the bottom of the well, the surrounding rock of the roadway at the tail roller section and the flat position section have different degrees of deformation under the excavation load. The basic theoretical breakthrough to solve this problem is to reveal the influence of excavation disturbance load on the deformation and failure of the main inclined shaft. The explicit finite difference program FLAC3D provides an effective method for studying this problem, which is especially suitable for simulating large deformation of the stope or roadway. In fact, the main cause of deformation and failure is the result of stress disturbance, and the stress conduction path of the main inclined shaft excavation disturbance is closely related to the construction method. The disturbance load characteristics of blasting construction are mainly a dynamic load. This paper does not consider the transmission of a dynamic shock wave, and only studies the deformation degree of the main inclined shaft from the results of disturbance after excavation, which has more practical significance for the deformation prediction and support optimization of the main inclined shaft excavated into roadways.

3. Numerical Model

The explicit finite difference program FLAC3D has been widely applied in the research and analysis of rock mechanics engineering [28,29,30] and can better simulate the mechanical behavior of failure, deformation or plastic flow in different lithologic strata when reaching the yield limit or strength limit, and analyze the progressive failure and instability, especially suitable for simulating large deformation. In this paper, aiming at the qualitative analysis of deformation caused by roadway excavation disturbance in the main inclined shaft of Yuxi Coal Mine, FLAC3D software is used to simulate and analyze the vertical and horizontal displacement evolution law under the influence of a bottom hole electromechanical chamber and head chamber excavation disturbance. The special flow chart illustrating the steps of numerical simulation and analysis is displayed in Figure 3.

3.1. Physical Model

Based on the geological conditions of the Yuxi coalfield, the numerical model of the main inclined shaft and related roadway is designed and constructed. Due to the excavation of the conveyor belt mechanical and electrical chamber and transportation belt roadway, in the main inclined shaft of Yuxi Coal Mine, there has twice occurred different degrees of surrounding rock deformation and failure. In order to investigate the influence of the excavation chamber and roadway on the displacement and stress redistribution of the surrounding rock of the main inclined shaft, according to the principle of proximity, the geological data of the YX-04 borehole in the mine are selected, and the regional strata (buried depth, 531–582 m) were selected as the engineering basis for the construction of the numerical model. The transportation belt roadway is excavated along the coal seam floor, and the Electromechanical chamber of the conveyor belt is perpendicular to the leveling section of the main inclined shaft. The size width, height and strike length of the model are 80 m, 50 m and 40 m, respectively. Midas GTX NX software is used to set the geometric parameters of the research model, and the parameters are matched with the actual size of the shaft and roadway in Yuxi Coal Mine. The formation model is constructed as shown in Figure 4.
The numerical model research object is mainly an inclined shaft, from +320 to +344 horizontal section, through 6 layers of rock. The lithology is mainly mudstone and sandy mudstone. The inclination angle of the wellbore is 16°, the azimuth angle is 258° and the length is 82 m (including the falling roadway). The wellbore is supported by the combined support of an anchor cable and bolt. Due to the engineering layout requirements (the tail of the conveyor belt is arranged at the levelling point), the electromechanical chamber is connected with the levelling roadway of the main inclined shaft, and the electromechanical chamber is arranged perpendicular to the levelling roadway in the horizontal direction, which is hosted in the similar rock strata. The lithology is sandy mudstone, and the chamber is supported by anchor cable. Due to the layout requirements of the lifting system, the head chamber of the transportation belt roadway is arranged in the coal seam perpendicular to the main inclined shaft in the horizontal direction. The floor of the coal seam in the bottom slab rock is excavated, and the support method is cable support. The shaft, roadway and chamber support models are shown in Figure 5.

3.2. Determination of Model Parameters

The formation parameters of wellbore, roadway and chamber are shown in Table 1. These parameters are input into the pre-established model of FLAC3D numerical simulation software in the form of commands. The main inclined shaft in the model passes through 6 layers of rock strata and is divided into 6 sections according to different rock strata. The transportation belt roadway occurs in 2 layers of rock strata, which is divided into 2 parts.
When using Midas GTX NX software to establish the model, the anchor cable and anchor node are marked out in advance, and the coordinate data are recorded. Edit these data and anchor, anchor parameters into the command input FLAC3D numerical simulation software pre-established model [31]. The parameters of the anchor are shown in Table 2.

3.3. Model Analysis and Design

The excavation before and after the electromechanical chamber and the head chamber resulted in the deformation and failure of surrounding rock in the horizontal section of the main inclined shaft from +320 to +344, mainly in the form of floor heave [32]. In order to analyze the deformation and failure mechanism of the surrounding rock of the main inclined shaft and the influence of the excavation of the two chambers on the main inclined shaft, the characteristic design analysis method is established based on the research purpose and model, as shown in Figure 6. Four sections perpendicular to the bottom plate of the model are selected for comparative analysis, i.e., X (10, 12), X (25, 27), X (50, 52) and X (67, 69). The stress and displacement under different excavation conditions are compared and analyzed along the section of the main inclined shaft.

4. Results

4.1. Effect of Stress Disturbance

4.1.1. Surrounding Rock Stress of Main Inclined Shaft without Excavation Disturbance

Figure 7, Figure 8, Figure 9 and Figure 10 show that the main inclined shaft is not disturbed by excavation, and the plastic zone range, anchor structural unit, stress distribution in the Z direction and Y direction of the four monitoring areas of the model section are taken. The analysis of the Mohr–Coulomb envelope shows that the structural unit of the plastic zone is formed by the influence of stable stress. For X (10, 12), the structural unit near the central axis of the roadway floor is the tensile force unit, and the structural units of the roadway side and roof are subject to shear force. Under the influence of shear force, a large-scale plastic zone appears in the strata under the roadway. The stress of the anchor cable is above 4.0 × 107 Pa, and the stress of the anchor bolt is about 2.0 × 107 Pa. The vertical principal stress of the roadway floor is 1.8232 × 105 Pa. The maximum shear stress of the two sides is 6.0 × 106 Pa, mostly concentrated in the bottom of the side.
For X (25, 27), the plastic zone range of roadway-surrounding rock increases, and the plastic zone affected by the shear force increases. The structural units near the central axis of the roadway floor are subjected to tensile tension and shear force, and the structural units of roadway side and roof are mostly subjected to shear force. The stress of the anchor cable is about 1.5 × 107 Pa, the stress range of the anchor bolt is about 2.5 × 106 Pa~2.5 × 107 Pa, and the maximum stress is the bottom angle anchor of the two roadways. The maximum vertical principal stress of roadway floor is 1.6896 × 105 Pa. The maximum shear stress of the two roadways is 6.0453 × 105 Pa, mostly concentrated in the surrounding rock of the roadway floor.
For X (50, 52), the plastic zone range of roadway-surrounding rock increases with the depth, and the maximum principal stress gradually changes from tensile stress to shear stress. The structural unit of roadway floor is affected by tensile stress and shear stress and does not reach a stable state. The structural unit of roadway side and roof is also affected by tensile stress and shear stress. The maximum stress of the anchor cable is 4.0067 × 107 Pa, and the stress range of the anchor rod is −1.1967 × 107 Pa~2.5 × 107 Pa, which is uniform. The vertical principal stress of the roadway floor is 2.8301 × 104 Pa, and the maximum vertical principal stress of the roadway side is −2.8947 × 107 Pa. From this analysis, it can be concluded that the floor heave is easily caused by vertical stress. The maximum shear stress of the two sides is −2.5340 × 107 Pa, and the shear stress decreases to −1.8942 × 105 Pa near the bottom corner of the two sides.
For X (67, 69), the plastic zone range of roadway-surrounding rock did not change significantly, and the maximum principal stress was shear stress. The stress form of roadway-surrounding rock structure unit changes, which is mainly caused by unstable stress. The maximum stress of the anchor cable is 3.7178 × 107 Pa, and the stress range of the anchor rod is −4.0161 × 106 Pa~2.5 × 107 Pa, which is uniform. The vertical principal stress of the roadway floor is 7.3724 × 105 Pa, and the maximum vertical principal stress of the roadway side is −3.5422 × 107 Pa. Under the combined action of the vertical stress of the two roadway sides and the free space movement on the floor of the roadway, it is easy for the floor heave phenomenon to occur in this section. The maximum shear stress of the two sides is −2.6063 × 107 Pa, and the shear stress decreases to −2.8282 × 105 Pa near the bottom corner of the two sides.

4.1.2. Surrounding Rock Stress of Main Inclined Shaft after Excavation of Electromechanical Chamber

Figure 11, Figure 12, Figure 13 and Figure 14 show the plastic zone range, bolt structural unit, stress distribution in Z direction and Y direction of the four monitoring areas of the section taken by the model after the excavation of the electromechanical chamber in the main inclined shaft. For X (10, 12), the plastic zone range of roadway-surrounding rock is formed by the combined action of stable tensile stress, stable shear stress and unstable shear stress. The roadway floor and the central axis are only affected by stable tensile stress, and the floor near the two sides is formed by shear stress and tensile stress. The maximum stress of the anchor cable is 7.5 × 106 Pa, the stress range of the anchor bolt is −5.7461 × 106 Pa~2.6177 × 107 Pa, and the maximum stress is the end of the bottom angle of the two sides. The vertical principal stress of the roadway floor is 8.7118 × 104 Pa, and the maximum vertical principal stress of the roadway side is −2.2340 × 107 Pa. The maximum shear stress on the roof of the roadway side is −2.4305 × 107 Pa, and the shear stress on the surrounding rock of the roadway decreases from the roof along the two layers to the roadway side and the floor. The minimum shear stress is −9.4713 × 105 Pa.
For X (25, 27), the plastic zone of roadway-surrounding rock is mainly formed by stable shear stress, and there is a plastic zone formed by a small amount of unstable shear stress and stable shear stress in the floor. The maximum stress of the anchor cable is 7.5 × 106 Pa, the stress range of the anchor bolt is −5.7641 × 105 Pa~2.6177 × 107 Pa, and the maximum stress is the end of the bottom angle of the two sides. The vertical principal stress of the roadway floor is 1.9796 × 105 Pa, and the maximum vertical principal stress near the head chamber of the belt roadway is −4.1689 × 107 Pa. The maximum shear stress of the roadway roof is −2.4516 × 107 Pa, and the minimum shear stress is −7.8189 × 105 Pa. The simulation calculation shows that the shear stress of the floor is much larger than that of the roadway roof.
For X (50, 52), the plastic zone of the surrounding rock of the first side of the tunnel and the floor is greatly affected by the unstable shear stress, but the plastic zone of the surrounding rock is still formed by the stable shear stress in a large range. The maximum stress of the anchor cable is 3.9828 × 107 Pa, and the stress range of the anchor rod is 5.0 × 106 Pa~3.9828 × 107 Pa. The maximum stress is the end of the anchor rod at the bottom corner of the two roadways. The vertical principal stress of the roadway floor is 2.6884 × 104 Pa, and the vertical stress range of the surrounding rock is −2.5 × 106 Pa~−3.8016 × 107 Pa. The local stress concentration area appeared on the roadway roof, with the maximum shear stress of −2.9983 × 107 Pa and the minimum shear stress of −2.0561 × 105 Pa. Compared with the unexcavated electromechanical chamber, the shear stress increases with an increment of 0.1619 × 105 Pa.
For X (67, 69), this interval is the main inclined shaft section near the electromechanical chamber. Due to the excavation of the electromechanical chamber, the surrounding rock of the main inclined shaft has a large area of plastic zone formed by unstable shear stress. The model grid is divided into 0.5 m/grid, so the unstable plastic zone of the two roadway sides has exceeded the support range, and the local unstable plastic zone of the roof reaches 3 m above the roof. The maximum stress of the anchor cable is 3.7409 × 107 Pa, and the stress range of the anchor bolt is −6.9376 × 106 Pa~3.25 × 107 Pa. The maximum stress is the rock bolt of the right side, which is the main stress concentration area. The vertical principal stress of roadway floor is 1.5523 × 105 Pa, and the vertical stress range of the surrounding rock is −2.5 × 106 Pa −3.7963 × 107 Pa. The maximum vertical principal stress appears in the surrounding rock of the right side of about 5 m. The local shear stress concentration area appeared on the roadway roof, with the maximum shear stress of −2.7255 × 107 Pa and the minimum shear stress of −2.5 × 106 Pa. The force of coal and rock mass transfers to the right due to the supporting effect.

4.1.3. Surrounding Rock Stress of Main Inclined Shaft after Excavation of Head Chamber in Belt Roadway

Figure 15, Figure 16, Figure 17 and Figure 18 show the plastic zone range, bolt structural unit and stress distribution in the Z direction and Y direction of the four monitoring areas of the cross-section taken by the model after the excavation of the head chamber of the belt roadway in the main inclined shaft. For X (10, 12), the range distribution of the stress-plastic zone of the surrounding rock in this interval has no significant change compared with that in the first two states. The maximum stress of the anchor cable is 4.3355 × 107 Pa, the stress range of the anchor cable is −1.0639 × 103 Pa~4.3355 × 107 Pa, and the stress is uniform and normal. The vertical principal stress of the roadway floor is 6.8863 × 104 Pa, and the vertical stress range of the surrounding rock is −2.0 × 106 Pa~−2.417 × 107 Pa. The maximum vertical principal stress appears in the surrounding rock of the two roadway sides. The range of the roadway roof and floor shear stress is −1.2357 × 106 Pa~−8.0 × 106 Pa.
For X (25, 27), the section is affected by the excavation of the head chamber of the belt roadway, resulting in a large area of the plastic zone formed by unstable shear stress in the surrounding rock of the main inclined shaft, except the right side. The unstable plastic zone of the rock mass under the bottom plate reaches more than 8 m. The maximum stress of the anchor cable is 2.9108 × 107 Pa, and the stress range of the anchor rod is −1.3686 × 106 Pa~2.9108 × 107 Pa. The bottom angle of the two roadways is the largest. The vertical principal stress of the roadway floor is 2.4889 × 105 Pa, and the vertical stress range of the surrounding rock is −2.5 × 106 Pa~−2.9042 × 107 Pa. The maximum vertical principal stress appears in the surrounding rock of the right side. The shear stress range of the roadway roof and floor is −1.0 × 106 Pa~−1.9113 × 107 Pa. The coal and rock mass in the bottom plate of the main inclined shaft moves to free space under the combined action of unstable vertical stress and shear stress, and bottom heave occurs.
For X (50, 52), the interval is affected by the excavation of the head chamber of the belt roadway. The plastic zone of the surrounding rock near the main inclined shaft falling roadway is mainly formed by the action of unstable tensile stress and shear stress, and the plastic zone range of the roof and floor is more than 8 m. The maximum stress of the anchor cable is 4.0958 × 107 Pa, and the stress range of the anchor bolt is 5.0 × 106 Pa~3.0 × 107 Pa. The vertical principal stress of the roadway floor is 3.3939 × 104 Pa, and the vertical stress range of the surrounding rock is −2.5 × 106 Pa~−3.8659 × 107 Pa. The maximum vertical principal stress occurs in the two groups of surrounding rock. The shear stress range of the roadway roof and floor is −1.8716 × 105 Pa~−3.0184 × 107 Pa.
For X (67, 69), this interval is the main inclined shaft section near the electromechanical chamber. Due to the excavation of the electromechanical chamber and the head chamber of the belt roadway, the plastic zone formed by the unstable shear stress of the surrounding rock of the main inclined shaft expands. The maximum stress of the anchor cable is 3.8204 × 107 Pa, and the stress range of the anchor bolt is −6.5227 × 106 Pa~3.25 × 107 Pa. The maximum stress is the rock bolt of the right side, which is the main stress concentration area. The vertical principal stress of the roadway floor is 2.4235 × 105 Pa, and the vertical stress range of the surrounding rock is −2.5 × 106 Pa~−4.0475 × 107 Pa. The maximum vertical principal stress appears in the surrounding rock of the right side of about 5 m. The local shear stress concentration area appeared on the roadway roof, with a maximum shear stress of −2.8584 × 107 Pa and a minimum shear stress of −2.5 × 106 Pa. The force of coal and rock mass transfers to the right due to the supporting effect.

4.2. Displacement Evolution Analysis

4.2.1. Surrounding Rock Displacement of Main Inclined Shaft without Excavation Disturbance

Figure 19 shows that the main inclined shaft is not disturbed by excavation, according to the Z direction and Y direction displacement evolution of the four monitoring areas of the section taken by the model. For X (10, 12), the maximum displacement of the main inclined shaft bottom plate is 0.19465 m. The maximum displacement of the two roadway sides is 0.22394 m and −0.23508 m, respectively, and the maximum deformation of roadway surrounding rock is determined to be about 0.2 m. For X (25, 27), the maximum displacement of the main inclined shaft floor is 0.31715 m. The maximum displacement of the two sides is 0.16951 m and −0.29814 m, respectively. For X (50, 52), the maximum displacement of the main inclined shaft bottom plate is 0.25871 m. The maximum displacement of the two roadways is 0.24599 m and −0.11255 m, respectively. The bottom angle of the two roadways is the deformation concentration area. For X (67, 69), the maximum displacement of the main inclined shaft floor is 0.35468 m. The maximum displacements of the two roadway sides are 0.35873 m and −0.11920 m, respectively. The bottom angle of the two roadway sides is the deformation concentration area. Because the surrounding rock of the main inclined shaft is six different lithology, the coal and rock mass moves in different degrees with the increase of buried depth, resulting in the increase of deformation of roadway with the increase of buried depth.

4.2.2. Surrounding Rock Displacement of Main Inclined Shaft after Excavation of Electromechanical Chamber

Figure 20 shows the displacement evolution in the Z direction and Y direction of the four monitoring areas of the cross-section taken according to the model after the excavation of the electromechanical chamber. For X (10, 12), the maximum displacement of the main inclined shaft bottom plate is 0.19483 m. The maximum displacement of the two sides is 0.22825 m and −0.23926 m, respectively. For X (25, 27), the maximum displacement of the main inclined shaft floor is 0.35279 m. The maximum displacement of the two sides is 0.19938 m and −0.33527 m, respectively. For X (50, 52), the maximum displacement of the main inclined shaft bottom plate is 0.42116 m. The maximum displacements of the two roadway sides are 0.48972 m and −0.34228 m, respectively. The bottom angle of the two roadway sides is the deformation concentration area. The excavation of the electromechanical chamber increases the deformation of the surrounding rock of the main inclined shaft, and the maximum floor heave is 0.42 m. For X (67, 69), the maximum displacement of the main inclined shaft floor is 0.64301 m. The maximum displacements of the two roadways are 0.84553 m and −0.63284 m, respectively. The bottom angle of the two roadways is the deformation concentration area. This section is at the intersection with the electromechanical chamber. It can be seen the maximum floor heave reaches 0.64 m, and the maximum convergence of the two roadway sides reaches 0.84 m. In summary, after the excavation of the mechanical and electrical chamber at the bottom of the well, a large area of surrounding rock deformation occurs in the falling roadway of the main inclined shaft, mainly due to the floor heave.

4.2.3. Surrounding Rock Displacement of Main Inclined Shaft after Excavation of Head Chamber of Belt Roadway

Figure 21 shows the displacement evolution in the Z direction and Y direction of the four monitoring areas of the cross-section taken according to the model after excavating the head chamber of the belt roadway. For X (10, 12), the maximum displacement of the main inclined shaft bottom plate is 0.18826 m. The maximum displacement of the two sides is 0.25924 m and −0.28947 m, respectively. For X (25, 27), the maximum displacement of the main inclined shaft floor is 0.43656 m. The maximum displacement of the two sides is 0.40807 m and −0.36228 m, respectively. The section is the intersection of belt roadway head chamber and main inclined shaft, and the deformation of surrounding rock is about 0.4 m. For X (50, 52), the maximum displacement of the main inclined shaft bottom plate is 0.45 m. The maximum displacement of the two roadways is 0.56614 m and −0.38844 m, respectively. The bottom angle of the two roadways is the deformation concentration area. The excavation of the head chamber of the belt roadway once again leads to the increase of the surrounding rock deformation of the main inclined shaft. For X (67, 69), the maximum displacement of the main inclined shaft floor is 0.65934 m. The maximum displacements of the two roadways are 0.87187 m and −0.64847 m, respectively. The bottom angle of the two roadways is the deformation concentration area. This section is at the intersection with the electromechanical chamber. It can be seen the maximum floor heave reaches 0.66 m, and the maximum convergence of the two roadway sides reaches 0.87 m. Overall, after the excavation of the head chamber of the belt roadway, the ground stress is redistributed due to the influence of mining disturbance, and the coal and rock mass move to free space. The cause is the intersection of the main inclined shaft and the head chamber to the main inclined shaft.

4.3. Analysis of Disturbance Influence Profile

Based on the above analysis, the stress and deformation of the surrounding rock mass are analyzed and studied along the main inclined shaft-cutting face under the influence of excavation disturbance, as shown in Figure 22. It mainly shows the evolution of stress and displacement in the Z direction of the section taken by the excavation model of the main inclined shaft (along the section of the main inclined shaft). Under the condition that the main inclined shaft is not affected by excavation disturbance, the stress range of floor-surrounding rock increases with the increase of buried depth. The maximum roof stress is 2.62 × 107 MPa, and the maximum floor stress is 8.37 × 106 MPa. After excavating the electromechanical chamber at the bottom of the well, the obvious bottom heave appears in the main inclined shaft section adjacent to the electromechanical chamber. In addition, after the excavation of the head chamber, the stress range of the surrounding rock of the main inclined shaft moves to the wellhead direction, and the stress increases. The maximum roof deformation is 0.247 m and the maximum floor deformation is 0.599 m without excavation disturbance. After the excavation of the electromechanical chamber, the obvious bottom heave phenomenon occurs in the horizontal section of the main inclined shaft. The deformation range of the floor near the chamber is 0.5–1.06 m, and the maximum deformation range of the roof is 0.862–1 m. The bottom drum range of the inclined shaft behind the chamber of the excavation head tends to increase. The literature points out that when the floor deformation reaches 1 m, the roadway section size will be greatly reduced, and high roadway maintenance costs will be incurred [33]. In summary, the serious deformation and failure area of the main inclined shaft mainly occurs near the intersection of the falling position, and the main causes are the deterioration of the surrounding rock fragmentation and the bottom heave caused by tensile stress. It is suggested that grouting reinforcement and the bottom anchor cable cooperative support method can be used for comprehensive treatment.

5. Discussion

With the continuous reduction of shallow coal resources, mining depth is deepening day by day during deep mining. Due to the superposition or redistribution of surrounding rock stress caused by the sequence of roadway construction times, some of the roadway groups arranged in soft rock are more prone to deformation and damage, which seriously affects the normal construction and production of the roadway. Therefore, how to control the deformation of the deep soft rock roadway and ensure the safety and stability of the surrounding rock is an urgent problem to be solved. The stress of roadway-surrounding rock and its support is complex, which is due to the complex geological occurrence conditions of the deep stratum, and the roadway may be affected by the construction of an adjacent roadway or chamber in the process of excavation. If the traditional mechanical theory is used to analyze and calculate the stress size and distribution law of roadway-surrounding rock and its supporting structure, it is very difficult. Based on the Yuxi coal mine with typical deep high stress, this paper studies the deformation evolution process and instability mechanism of the surrounding rock of the deep main inclined shaft by using the FLAC3D numerical simulation method. Compared to traditional mechanical theoretical analysis methods, the numerical simulation method is able to grasp the deformation of the deep tunnel surrounding rock and the support structure force situation in time to optimize the roadway support plan.
Although many scholars have done a lot of research on the deformation and failure mechanism of roadway-surrounding rock, the disasters caused by roadway instability are still very common in deep mining engineering. The current research mainly focuses on the formation of the stress adjustment stage in the process of deep tunnel excavation. There are few studies on the law of stress distribution in the process of tunnel excavation. The factors affecting the rock stability around the roadway are complex and vary according to the geological conditions of different coal mines. The existing research data are not necessarily applicable to the Yuxi coal mine. Moreover, due to the excavation of the mechanical and electrical chamber of the conveyor belt and transportation belt roadway, the surrounding rock deformation and failure of different degrees occurred in the main inclined shaft of the Yuxi coal mine twice. According to the actual situation of this site, this paper simulates the stress state of the surrounding rock of the main inclined shaft after excavating the electromechanical chamber and the machine head chamber of the belt roadway, studies the failure form variable of the surrounding rock and summarizes the redistribution characteristics of the stress field of the main inclined shaft after being disturbed by mining. Compared with the previous work done by scholars, the work of this paper is more targeted, the actual engineering size is restored 1:1, and the influence of two different excavation chambers on the surrounding rock of the roadway is simulated. The research content can provide some guidance for the deep excavation of the Yuxi coal mine and similar projects in other areas. Indeed, the research work of this paper is only carried out according to the geological conditions of the Yuxi coal mine, and the failure characteristics and modes of surrounding rock under more complex mine conditions need to be further discussed.

6. Conclusions

This paper takes the deformation and failure of surrounding rock of the main inclined shaft in Yuxi Coal Mine as the engineering background and uses Midas GTX NX to complete the 1:1 engineering modeling. Then, the explicit finite difference program (FLAC3D) is used for numerical calculation of the model, and the simulation results are analyzed in detail. The following conclusions can be drawn:
(1) With the increase of buried depth, coal and rock mass induce different degrees of movement, so that the deformation variable of the roadway increases with the increase of the buried depth. The surrounding rock stress of the main inclined shaft decreases under the influence of excavation disturbance. For monitoring area X (10, 12), after the excavation of the electromechanical chamber, the vertical principal stress of the roadway floor decreases from 1.8232 × 105 Pa to 8.7118 × 104 Pa; after excavating the head chamber of the belt roadway, the change value is reduced from 8.7118 × 104 Pa to 6.8863 × 104 Pa. The coal and rock mass in the bottom plate of main inclined shaft moves to free space under the combined action of unstable vertical stress and shear stress, resulting in the occurrence of floor heave.
(2) The displacement evolution laws of different areas of the main inclined shaft are different after being disturbed by excavation. The maximum displacement of the bottom plate in the monitoring area X (67, 69) is sensitive to the excavation disturbance and is most affected by the excavation disturbance. The maximum displacement of the floor of the excavation electromechanical chamber and the head chamber of the belt roadway increased from 0.35468 m to 0.64301 m and 0.65934 m, respectively, nearly doubling. At that time, a large area of surrounding rock deformation occurs in the main inclined shaft roadway, and the bottom heave is the main phenomenon.
(3) Under the influence of the excavation disturbance, the scope of the inclined shaft bottom heave after excavating the machine head chamber tends to increase. The maximum floor deformation can reach 1.06 M, exceeding the limit of 1 m. The fluctuation range is wide, which will seriously restrict the production efficiency of coal mine. The serious deformation and failure area of the main inclined shaft mainly occurs near the intersection of the falling position. The main causes are the deterioration of the surrounding rock fragmentation and the bottom heave caused by tensile stress. The damage to the main inclined shaft can be reduced by grouting reinforcement at this position and the cooperative support of a floor anchor cable.

Author Contributions

Conceptualization, Y.Q. and H.X.; Data curation, F.W., H.X. and F.Z.; Investigation, F.W., H.X., F.Z. and X.C.; Methodology, F.W.; Project administration, Y.Q.; Resources, F.W.; Supervision, Y.Q.; Visualization, Y.Q., H.X., F.Z. and X.C.; Writing—Original draft, F.W.; Writing—Review & editing, H.X. All authors have read and agreed to the published version of the manuscript.

Funding

This research was funded by [National Natural Science Foundation of China] grant number [51874315; 52074303] and [Fundamental Research Funds for the Central Universities] grant number [2021YJSAQ24].

Institutional Review Board Statement

Not applicable.

Informed Consent Statement

Not applicable.

Data Availability Statement

Not applicable.

Conflicts of Interest

The authors declare no conflict of interest.

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Applsci 12 05531 g001 550
Figure 1. Geographical location of mine.
Figure 1. Geographical location of mine.
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Figure 2. Geographical location of the mine.
Figure 2. Geographical location of the mine.
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Figure 3. Flow chart illustrating the steps of the numerical analysis.
Figure 3. Flow chart illustrating the steps of the numerical analysis.
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Figure 4. Actual engineering geological physical model.
Figure 4. Actual engineering geological physical model.
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Figure 5. Supporting model of shaft, roadway and chamber in practical engineering.
Figure 5. Supporting model of shaft, roadway and chamber in practical engineering.
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Figure 6. Model design and measuring point arrangement.
Figure 6. Model design and measuring point arrangement.
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Figure 7. Plastic zone range of the four monitoring areas without excavation disturbance.
Figure 7. Plastic zone range of the four monitoring areas without excavation disturbance.
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Figure 8. Cable force of the four monitoring areas without excavation disturbance.
Figure 8. Cable force of the four monitoring areas without excavation disturbance.
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Figure 9. Z−direction stress distribution of the four monitoring areas without excavation disturbance.
Figure 9. Z−direction stress distribution of the four monitoring areas without excavation disturbance.
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Figure 10. Y−direction stress distribution of the four monitoring areas without excavation disturbance.
Figure 10. Y−direction stress distribution of the four monitoring areas without excavation disturbance.
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Figure 11. Plastic zone range of the four monitoring areas after excavation of electromechanical chamber.
Figure 11. Plastic zone range of the four monitoring areas after excavation of electromechanical chamber.
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Figure 12. Cable force of the four monitoring areas after excavation of electromechanical chamber.
Figure 12. Cable force of the four monitoring areas after excavation of electromechanical chamber.
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Figure 13. Z−direction stress distribution of the four monitoring areas after excavation of electromechanical chamber.
Figure 13. Z−direction stress distribution of the four monitoring areas after excavation of electromechanical chamber.
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Figure 14. Y−direction stress distribution of the four monitoring areas after excavation of electromechanical chamber.
Figure 14. Y−direction stress distribution of the four monitoring areas after excavation of electromechanical chamber.
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Figure 15. Plastic zone range of the four monitoring areas after excavation of head chamber in belt roadway.
Figure 15. Plastic zone range of the four monitoring areas after excavation of head chamber in belt roadway.
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Figure 16. Cable force of the four monitoring areas after excavation of head chamber in belt roadway.
Figure 16. Cable force of the four monitoring areas after excavation of head chamber in belt roadway.
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Figure 17. Z−direction stress distribution of the four monitoring areas after excavation of head chamber in belt roadway.
Figure 17. Z−direction stress distribution of the four monitoring areas after excavation of head chamber in belt roadway.
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Figure 18. Y−direction stress distribution of the four monitoring areas after excavation of head chamber in belt roadway.
Figure 18. Y−direction stress distribution of the four monitoring areas after excavation of head chamber in belt roadway.
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Figure 19. Main inclined shaft and supporting state without excavation disturbance.
Figure 19. Main inclined shaft and supporting state without excavation disturbance.
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Figure 20. Main inclined shaft and supporting state after excavation of electromechanical chamber.
Figure 20. Main inclined shaft and supporting state after excavation of electromechanical chamber.
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Figure 21. Supporting state after excavation of belt roadway head chamber.
Figure 21. Supporting state after excavation of belt roadway head chamber.
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Figure 22. Stress and displacement profile of surrounding rock of main inclined shaft disturbed by excavation.
Figure 22. Stress and displacement profile of surrounding rock of main inclined shaft disturbed by excavation.
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Table
Table 1. Basic parameters of rock strata in shaft, roadway and chamber roof and floor.
Table 1. Basic parameters of rock strata in shaft, roadway and chamber roof and floor.
Strat
Graphic Number		HorizonLithologyLayer Thickness (m)Volume Modulus (GPa)Shear Modulus (GPa)Tensile Strength (MPa)Cohesion (MPa)Internal Friction Angle (°)Bulk
Density (KN·m−3)
GS-No. 01Overlying strataSandy mudstone2.002.811.031.410.72225.426.1
GS-No. 02Overlying strataMudstone2.152.811.031.410.72225.426.1
GS-No. 03Overlying strataMedium sandstone5.504.072.581.821.41027.427.9
GS-No. 04Overlying strataMudstone3.852.811.031.410.72225.426.1
GS-No. 05Overlying strataFine-grained sandstone1.303.872.131.641.26026.326.5
GS-No. 06Overlying strataSandy mudstone1.452.811.031.410.72225.426.1
GS-No. 07Overlying strataSiltstone1.902.821.051.430.73626.427.3
GS-No. 08Main inclined shaftMudstone3.552.811.031.410.72225.426.1
GS-No. 09Main inclined shaftSandy mudstone1.122.811.031.410.72225.426.1
GS-No. 10Main inclined shaftCoal seam5.855.200.380.642.03026.013.0
Belt alleyCoal seam5.855.200.380.642.03026.013.0
GS-No. 11Main inclined shaftMudstone0.572.811.031.410.72225.426.1
Belt alleyMudstone0.572.811.031.410.72225.426.1
GS-No. 12Main inclined shaftSiltstone4.512.821.051.430.73626.427.3
GS-No. 13Main inclined shaftSandy mudstone7.752.811.031.410.72225.426.1
GS-No. 14Underlying strataLimestone0.802.301.942.603.35038.727.4
GS-No. 15Underlying strataMudstone7.702.811.031.410.72225.426.1
Table
Table 2. Basic physical and mechanical parameters of anchor cable and anchor in main inclined shaft.
Table 2. Basic physical and mechanical parameters of anchor cable and anchor in main inclined shaft.
NameLength (mm)Exposure (mm)Spacing (mm)Line Spacing (mm)Cross
Sectional Area (mm2)		Cohesion (GPa)Anchoring Force (GPa)Elastic Modulus (GPa)Tensile Strength (MPa)
Anchor cable73003001200240011040201.3
Bolt24003001200100011022010
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Wu, F.; Qin, Y.; Xu, H.; Zhang, F.; Chu, X. Numerical Simulation of Deformation and Failure Mechanism of Main Inclined Shaft in Yuxi Coal Mine, China. Appl. Sci. 2022, 12, 5531. https://doi.org/10.3390/app12115531

AMA Style

Wu F, Qin Y, Xu H, Zhang F, Chu X. Numerical Simulation of Deformation and Failure Mechanism of Main Inclined Shaft in Yuxi Coal Mine, China. Applied Sciences. 2022; 12(11):5531. https://doi.org/10.3390/app12115531

Chicago/Turabian Style

Wu, Fan, Yueping Qin, Hao Xu, Fengjie Zhang, and Xiangyu Chu. 2022. "Numerical Simulation of Deformation and Failure Mechanism of Main Inclined Shaft in Yuxi Coal Mine, China" Applied Sciences 12, no. 11: 5531. https://doi.org/10.3390/app12115531

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