1 Introduction

At present, in order to increase the resource recovery rate, coal area are often abandoned that are difficult to recover due to factors, such as the boundary of wells, geological structural belts, coal seam change belts, and protection pillars. So the arrangement of the “Knife handle” type working face is usually carried out. Sometimes, in order to avoid taking over the tension, we use adopting the jump-picking method so that an isolated coal face is formed inevitably in the production of coal mines. In terms of island working face, due to the influence of the goafs on both sides of the working face, the stress in the face coal body is superimposed, the peak value is increased, and stress-concentration area is expanded, that resulting in a large compressive strength at the top the face and a strong mine pressure. Especially in the stress concentration area, there is the possibility of rock burst (Li 2017; Sun 2008; Li et al. 2017; Lu et al. 2016; Ding 2014; Li et al. 2015a, b). With the increasing mining depth and mining intensity, the deep mining causes a series of problems and dynamic hazard is becoming more likely to happen, especially rock burst (Guo et al. 2017; Wen et al. 2016). In mining process, In order to understand the mechanism of rockburst occurrence, a large number of studies have been conducted to obtain a comprehensive understanding of rock burst. Specifically, the focus of these studies have been the mechanism of rock burst, overburden structure movement, abutment stress distribution (Jiang et al. 2012; Li et al. 2016a, b) and assessment of rock burst risk (Li et al. 2015a, b; Zhu et al. 2015; Christopher and Michael 2016). Many scholars acknowledged that rockbursts are a sudden and violent coal mass failure phenomenon, which has a relationship with the instantaneous release of strain energy (Wang et al. 2017; Cook 1965; Brown 1998; Salamon 1984; Singh 1988; Tan and Zhang 2012). The special and complex overburden structure of the of the “Knife handle” type working face which make it easy to produce high stress concentration in the changing area. It is easy to cause a sudden instability of the coal body due to external disturbances, but few studies have been done about the mechanical characteristics of the overburden fracture of the blade stope, the relationships between overburden movement and bearing pressure, and the rock mass migration and mine pressure (Jiang 2006; Gao and Wang 2013). Nowadays, there are few studies on rock burst in the lower seam of the “knife handle” type isolated coal face, and the theory is not so mature. There is no evidence of actual observations in this aspect. In addition to, the mechanical properties of coal and rock in different areas, mining geological conditions, and coal and rock properties are quite different, and the factors of inducing are different. As a result, the mechanism of rock burst occurrence is complicated and changeable. Therefore, we still need to go further study of the mechanism and control technology of rock burst occurrence (Li et al. 2016a, b; Han et al. 2003; Yu et al. 2013). Due to surrounding rock of mining gateway isolated coal Face with the characteristics of large deformation, hard support and stress concentration. In this paper, the main influencing factors of the rock burst on the slicing “Knife handle” type isolated island hauglage gateway are theoretically analyzed based on the geological conditions of the 1304 working face in a mine. Then the FLAC 3D5.0 is used to establish numerical model for the main influencing factors to simulate and analyze the criticality during the excavation and mining. Taking precautions against dangerous areas in advance in order to reduce the rock burst on Mining Roadway utmostly. It has guiding significance for the excavation and maintenance of mining roadway in working face with similar geological conditions.

2 Project Overview

1304 is a typical unequal length working face with a length of 57.1–160.7 m. The direction of the working face is 899 m long, the facet whose direction is 57.1 m is 110 m long, large-face joined length of 160.7 m goes to 789 m long. There are 1304 upper, 1302 upper and 1316 upper goafs areas above the working face 1304, and the remaining isolated coal pillars are formed between goafs. There are 46 m long isolated coal pillars between rail roadway of 1304 upper to 1316 upper. The rail roadway of 1304 coal face are under the isolated coal pillar, and haulage gateway is under the goaf of 1304. There are 5 m long coal pillars between track lanes of 1304–1316, and a section of coal pillar with 14.3 m between the 1304 working face haulage gateway and 1302 goaf under 1304 goaf; the west side of the 1304 working face is 1316 face goaf, and the width of the coal pillar in the two working face section is 5 m. The coal seam thickness of working face is 6.8–10.6 m, the average value is 7.5 m, upper seam has been 3.5 m and lower layer has been 4 m. Surrounding environments and mining conditions are more complex in the working face 1304, the deformation and stress concentration of the surrounding rock in the 1304 roadway are large. Therefore, the main influencing factors on the rock burst hazard of the mining roadway on 1304 working face are studied, which is of guiding significance to the mining of the “knife handle” type island working face in the lower seam with similar geological conditions. The layout of the roadway in the working face is shown in Fig. 1.

Fig. 1
figure 1

Layout of roadway in working face. a Plan graph of roadway layout in working face. b Sectional view of roadway layout in working face

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3 Analysis on the Main Impact Factors of Rock Burst Hazard in Rail Roadways

3.1 Rock Burst Hazard Analysis of Isolated Coal Pillar Above 1304 Working Face

The original stress balance is broken after mining on both sides of the working face, and the weight of the overlying rock will be transferred to the coal and rock mass around the goaf, thus the bearing pressure area is formed, which are three kinds of complex stress separation: front abutment pressure, the lateral support pressure and the bearing pressure in the goaf. These three kinds of complex stresses are: (1) the bearing stress of rock stratum failure and movement formed in the range of mining influence; (2) lateral overlying strata fracture in both sides of goaf, fractured rock formation transfer rock beams, self weight of marginal strata and lateral supporting stress formed by residual tectonic stress in the upper layer mining area; (3) The gravity stress field formed by overlying rocks on isolated coal pillars (Chen et al. 2016; Li et al. 2010). The front support pressure in front of the working face is superimposed on the lateral support pressure formed by the upper stratified goaf, and then is coupled with the self gravity stress field formed by the overlying rock of the isolated island coal pillars. Under the superposition of these three complex stresses, the coal pillar increases the degree of stress concentration and the high strength elastic deformation energy. In addition, the solid coefficient f of coal body in 1304 working face is more than 1.5, the coal rock mass is influenced greatly by the tectonic movement and mining disturbance. Under the coupling of the complex stress field, the isolated coal pillar is prone to dynamic instability, which leads to the occurrence of the rock burst (Wang et al. 2013). The 1304 working face rail roadway is under the isolated island coal pillar, and the outer error is arranged in 1304 goaf, while the stress concentration is produced in the isolated island coal pillar, and it will be transferred to the track roadway below the coal pillar. The rail roadway is protected by 5 m long coal pillar uses the method of driving roadway along gob, which is affected by the direct top caving and the lateral basic top key, resulting in a great stress concentration. Most of the pillar coal pillars are in the plastic state, and the bearing capacity is very small or near the non load-bearing capacity. The maximum stress concentration occurs, due to immediate roof caving of the upper section and the rotation of the key blocks on the lateral key roof. Most section of the protective coal pillar enters the plastic state, bearing capacity is very small or almost no. The concentrated stress produced by three complex stresses is transferred to the surrounding solid coal, which leads to the large bearing capacity of the solid coal, the solid coefficient of solid coal is about 1.5, and it is easy to accumulate energy. Under the influence of rock burst factors, it is easy to cause energy release suddenly, resulting in the occurrence of rock burst.

3.2 Rock Burst Risk Analysis of Rail Haulage Gateway During the “Knife Handle” Isolated Face is Changing

For “knife handle” coal face, there are three development stages along the direction of the working face, including small face stope, face variable stope and large face stope, and three transformation processes of slow pressure static mining area, different pressure variable mining area and high pressure mining area (Wang 2015), are experienced in each stage of Mining overlying rock. According to the occurrence regularity and movement characteristics of strata in different mining stages, the spatial structure of overburden strata can be divided into three types: high pressure structure, catastrophe structure and slow pressure structure. The overlying rock structure model of “three fields, three districts and three structures” is obtained. As shown in Fig. 2.

Fig. 2
figure 2

“Knife handle” type coal face structure model

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As shown in Fig. 2, The second periodic weighting of small face stope are pushed forward to the position of variable face stope, and the key layer rock stratum breaks down. At this time, the roof rock layer is broken down in large area, a number of cantilever beams are formed in the variable face stope. The length of cantilever beam will become longer in a more high-order position. With the combined action of roof weight, mining stress and advanced abutment pressure, the high stress concentration area of coal and rock mass is formed. The “rock-coal” strength theory pointed out that when the rock pressure is greater than the comprehensive strength of the “rock and coal” mechanical system, the coal-rock mass break down and lose its stability so that rock burst occurs (Tan 2011). According to the theory of coal-rock clamping, the limit equilibrium conditions of “rock-coal” mechanical system are obtained and the ultimate compressive stress formula for the “rock and coal” system is deducted (Zhang et al. 2015)

$$P_{K} = \left( {\sigma_{C} + \tau_{0} \tan \varphi } \right)\left[ {\exp \left( {\tan \varphi \frac{x}{M}} \right) - 1} \right] + \sigma_{C}$$
(1)

where σc is single axis pressure resistant intensity, σc = 7.63 MPa; τ0 is “Surface-coal” interface cohesion, τ0 = 2.0 MPa; φ is “Surface-coal” interface friction angle, φ = 30°; x is distance of calculation point from coal wall, x = 10 m; M is Mining height, M = 7.5 m.

According to the formula of ultimate compressive stress of “rock-coal” system, the ultimate compressive stress of coal at 10 m in front of the coal wall of 1304 face is calculated to be 17.9 MPa. With the working face advancing gradually, mining disturbances increase and the stress peak shifts to the deep part of the coal and rock mass. When the small face is mined for 100 m, there is a variable-face stope in the different-pressure transformation area 10 m ahead of the coal wall in working face. According to the on-site stress monitoring, it shows that support pressure in front of the coal wall 10 m long is about 41.0 MPa, and the front abutment pressure is about 2.3 times the ultimate compressive stress of the coal rock mass. With such high pressure conditions, the roadway is susceptible to rock burst due to mining disturbances or other shock hazards.

4 Distribution of Abutment Pressure and Displacement Deformation in Roadway with “Knife-Knife” Isolated Coal Face

4.1 Model Dimensions and Mechanical Parameters

The peak value of the front abutment pressure for the driving roadway under the isolated coal pillar is larger than that of the conventional driving roadway, and the stress concentration coefficient is about twice that of the conventional case (Zhang et al. 2014). The stress concentration at the face of mining disturbance is even greater. Seriously, it has a great impact on the stability of rail roadway. This paper uses FLAC3D5.0 to conduct numerical tests on isolated coal pillars and rail roadway under coal pillars. A numerical model was established, length × width × height = 700 m × 600 m × 111 m, a total of 546,000 grids were divided, the sides and bottom of the model were fixed, the vertical uniform load of 7.5 MPa was applied on the top to simulate the rock weight of the overburden, simulation using Mohr–Coulomb yield criteria. The rock mechanics parameters are shown in Table 1.

Table 1 Rock mechanical parameters
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4.2 Analysis of Simulation Results

  1. 1.

    “Knife-handle” type isolated coal face support pressure and plastic zone distribution.

    On the coal face of 1316 upper and 1304 upper, there is 46 m long isolated coal pillar between two working faces. The lateral support pressure of two adjacent goafs acts on the isolated coal pillar to form a superposition effect of the support pressure, as shown in Fig. 3. The two sides of the support pressure formed two stress concentration areas on the edge of the coal pillar, the lateral support pressure peak value was about 35.0 MPa near 1316 upper working face, the original rock stress was 11.5 MPa, and the stress concentration factor was 3; the peak side abutment pressure is about 37.5 MPa near the 1304 upper working surface. In the upper strata goaf, and the stress concentration factor is 3.2, as shown in Fig. 4 It can be seen that the stress concentration of the layered goaf on working face of the isolated coal pillar near 1304 is slightly higher than that of the goaf area in 1316 coal face. The support pressure in the middle of the isolated pillar is 23.2 MPa, the stress concentration factor is 2 and higher than original rock stress, therefore, the stress on both sides of the isolated pillar has been superimposed on the middle of it, so that the isolated coal pillar is under pressure.

    Fig. 3
    figure 3

    Abutment pillar pressure

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    Fig. 4
    figure 4

    Abutment pressure distribution

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    Under the influence of the side abutment pressure of adjacent goafs, the isolated coal pillar is under pressure. Under the effect of superimposed bearing pressure, all plastic deformation of isolated coal pillar occurs, as shown in Fig. 5. Considering the formation of stress concentration in the isolated coal pillar, that means the residual strength of coal pillar has a certain bearing capacity. Therefore, the isolated coal pillar form a yield pillar, which can support the load imposed by the overlying strata. However, due to the full plasticity damage of the isolated coal pillar. The rail roadway under the isolated pillar should be reinforced to prevent and control the large deformation of the roadway when lower seam face 1304 is mined.

    Fig. 5
    figure 5

    The plastic zone of isolated coal pillar

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  2. 2.

    Evolution law of stress field and displacement field during track drifting.

    As shown in Fig. 6a, on the side of isolated coal pillar, the side abutment pressure is increased to 36.0 MPa and the stress concentration factor is 3.1. Compared with the undigging rail roadway, the stress concentration of the isolated coal pillar only increases slightly, mainly because of the roadway has a few influence on the excavation of surrounding rock, and the supporting pressure caused by the coal body increases less. It can be seen that there is a greater influence on the surrounding rock stress of rail roadway is the side abutment pressure generated by the adjacent goaf.

    Fig. 6
    figure 6

    The stress distribution and deformation of surrounding rock during rail roadway. a Roadway rock stress distribution. b Two sides displacement. c Top floor displacement

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    Under the effect of lateral support pressure, the subsidence of the rail roadway roof is mainly caused by the sinking of the roof in the adjacent goaf. For the two sides of the roadway, the deformation of the section coal pillar side is 175.0 mm, and the deformation of isolated coal pillar side is 76.0 mm, as shown in Fig. 6b, c. It can be seen that the deformation of rail roadway mainly consists of its two sides. Due to the full plasticity damage of the isolated coal pillar, compression deformation occurs when coal is compressed and it will play a role of pressure relief. Then the stress on the bottom plate is reduced and the floor heave of the roadway floor is reduced.

  3. 3.

    The distribution of abutment pressure and deformation laws of rail roadway during working face mining.

When lower seam “knife-handle” type isolated coal face 1304 is mined, the isolated coal pillar is in a open state on the three sides. The lateral abutment pressure of the gob on the two adjacent working faces and the leading bearing pressure of working face 1304 formed superimposed stress on isolated pillar, and the degree of stress concentration is higher, as shown in Fig. 7. When the working face is mined for 100 m long, ahead of the coal face of the working face 10 m long is a variable-faced stope in the different-pressure zone. The maximum abutment pressure at both edges of the working face is about 44.0 MPa, and the stress concentration factor reaches 3.8. The front abutment pressure has a severe impact. As the distance from the coal face of working face is farther and farther away from the variable pressure zone, although the peak value of support pressure is lower, the concentration is still higher.

Fig. 7
figure 7

cloud chart distribution of supporting pressure of isolated coal pillar during working face 1304 mining. a Distance from the face is 10 m. b Distance from the face is 20 m. c Distance from the face is 30 m. d Distance from the face is 40 m

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During lower seam “knife-handle” type working face 1304 is mining, the three-direction support pressure acts on the coal pillar, resulting in the coal pillar being under high pressure state. Due to the small strength, the coal body undergoes large deformation under the superimposed stress, as shown in Fig. 8. There is a variable face stope in the variable pressure and different mining area in front of the coal wall 10 m long on working face when the working face is mined for 100 m long. The deformation of the section coal pillar side in rail roadway is 446.0 mm, and the deformation of isolated coal pillar side is 265.0 mm. This is a zone with a drastically-impacted effect. With the change of the stope, the deformation of the roadway is reduced. However, when the coal wall is 40 m away from the coal face, the deformation of the section coal pillar side is 340.0 mm, and the deformation of the isolated pillar side is 160.0 mm, the total deformation of the two sides is still large.

Fig. 8
figure 8

Cloud map of the deformation of roadway during 1304 working face mining. a Distance from the face is 10 m. b Distance from the face is 20 m. c Distance from the face is 30 m. d Distance from the face is 40 m

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5 On-Site Monitoring

5.1 Front Abutment Pressure Observation

The range of observation that can be affected by front abutment pressure is 70 m, and 28 measurement points are arranged. When working face mining 100 m, four measuring points at 10 m (1#), 30 m (2#), 50 m (3#), and 70 m (4#) from the coal face are observed. One single prop is supported at each point and an infrared transmission pressure recorder is installed. The recorder recorded the change of individual prop load of each measuring point with the advancing distance of the working face, and analyzed four measuring points respectively, as shown in Fig. 9.

Fig. 9
figure 9

Changing curve of advance load prop. a 1#. b 2#. c 3#. d 4#

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From Fig. 9, it shows that the load changes of several measuring points are basically similar. With the coal face moving on, the resistance of the props of each measuring point has experienced a process from small change to obvious increase, which objectively reflects the change in front abutment pressure. The area where the prop resistance significantly increases is the plastic zone of the front roadway that is the zone with a drastically advanced impact. The area where the prop resistance starts to rise slightly with the advancing of the working face is the affected obviously by the front abutment pressure. The resistance of the pillar at 1# and 2# measuring points decrease suddenly when they are about 10.5 m away from the working face and they begin to rise significantly at around 9 m. The 3# and 4# measuring points are installed respectively at 50 m and 70 m away from the working face. The amplitude of support resistance increasing at a distance of about 41 m from the working face becomes larger. At this time, the pressure of the individual struts is larger. The resistance of the pillar suddenly increases again when the strut is 19.5 m away from the working face, the amplitude is larger and it will reach to the peak at around 11 m long of coal face.

5.2 Observation of Displacement Deformation of Two Gangs of Roadway

There are four surface displacement measuring stations in the rail roadway, namely 1#, 2#, 3#, and 4#, When working face mining 100 m, four measuring points at 10 m (1#), 30 m (2#), 50 m (3#), and 70 m (4#) from the coal face are observed, and the average distance between the stations is about 20 m. Each station uses a mine rifle to measure the distance of two sides in rail roadway. The changing curve of the 1#–4# shifts with advancing face is shown in Fig. 10.

Fig. 10
figure 10

Displacement deformation curves of two pairs of roadway. a 1#. b 2#. c 3#. d 4#

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From the four measuring points in Fig. 10, it can be seen that the deformation of the side near section coal pillar is far greater than the deformation of the side on isolated coal pillar, the amplification of deformation of two sides on coal pillars increase at the distance from the working face around 40 m. When the pillar is about 19.5 m away from the working face, the resistance of the pillar increases again suddenly, the amplification is larger, and it will reach a peak value of about 722.0 mm at about 11 m from the working face.

According to the analysis of the forward pillar load observation data and displacement deformation of the roadway, the preceding influence range is about 41 m, and the obvious influence range is about 20 m. The plastic zone appears in the top plate in the range 21 m ahead of the working face because of the effect of advance support pressure. Through the observation data analysis of forward support pillar load, the preceding influence range is about 41 m, and the obvious influence range is about 20 m. The plastic zone appears in the top plate within the range of 21 m at the front of the working face because of the effect of the front abutment pressure. With the advancing of working face, the support pressure increases continuously and reaches its maximum value when the rock beam breaks. The leading peak stress of the working face is about 11 m away from the working face, and the fracture position is about 10.5 m from the working face. After the failure of the rock beam, the support bearing pressure distribution of the fault line is clearly divided into two parts, namely the internal stress zone and the external stress zone. The extent of the plastic zone in the coal wall shifts to the two stress fields inside and outside and gradually stabilizes.

6 Conclusion

  1. 1.

    Through the theoretical analysis, it is found that the layout of the “knife-handle” type isolated working face and the isolated coal pillar left by the upper layer mining have a great influence on the rock burst risk of the rail roadway. The coal pillar under the effect of three kinds of complex stress superposition, which increases the degree of stress concentration. Most part of the coal pillar enter the plastic state and the bearing capacity is extremely small. The coal pillar transmits the concentrated stress generated by the three kinds of complex stresses to the solid coal around the roadway, which leads to the greater stress concentration and displacement deformation of the rail roadway. The variable face stope is located at the breaking point of the key strata. At this time, the roof rock layer is broken down in large area, a number of cantilever beams are formed in the variable face stope, and the cantilever is longer in a higher position. Under the combined action of dead weight of the top rock strata, mining stress and front abutment pressure of the goaf, the coal face with high stress concentrates into a high stress concentration zone. The high stress concentration area is formed of coal and mass in the variable face stope.

  2. 2.

    The analysis of the numerical model shows the deformation of the roadway side is larger and roof and floor is smaller. The displacement of the side pillars of the section coal pillars is larger than that of the side pillars of the island pillars. During the working face mining, the stress concentration is higher in front of the working face. There will be a variable face stope in the variable pressure zone in front of the coal face 10 m at the working face, when the small face stope is pushed forward by 100 m. The supporting pressure on the two edges of the working face reaches its extreme value. The stress concentration degree is very high, the deformation of section coal pillar side and the isolated pillar side is extremely large in the rail roadway.

  3. 3.

    It can be known from the on-site advance bearing pressure and roadway displacement deformation monitoring that the leading influence range is about 40.5 m, and the obvious influence range is 20 m. The internal and external stress fields are formed when the basic roof is pressed and the roof rock breaks. The boundary of the stress field is about 9 m in front of the working face. Therefore, when the rail roadway is within the range of front abutment pressure, the deformation of the roadway substructure increases, so that supporting strength needs to be increased properly. At the same time, it is necessary to adopt an effective method to reduce the stress of surrounding rock at the variable face and to promote the transfer of the stress peak move to the direction that is away from the deep surrounding rock around the roadway, and the stress rise area is avoided, so that the roadway is in the stress reduction area, so as to achieve the purpose of maintaining the stability of the roadway.