Numerical Simulation of the Stress Field in Repeated Mining of Coal Seams Based on <i>In Situ</i> Stress Measurement

Yang, Donghui; Yang, Bohu; Lv, Zhaoheng; Li, Yongming; Cheng, Hongming; Ning, Zhangxuan

doi:https://doi.org/10.1155/2021/3588868

Mathematical Problems in Engineering

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Volume 2021 | Article ID 3588868 | https://doi.org/10.1155/2021/3588868

Numerical Simulation of the Stress Field in Repeated Mining of Coal Seams Based on In Situ Stress Measurement

Donghui Yang,¹Bohu Yang,¹Zhaoheng Lv,¹Yongming Li,¹Hongming Cheng,¹and Zhangxuan Ning¹

Academic Editor: Qiuhong Wu

Received09 Aug 2021

Accepted27 Sept 2021

Published09 Nov 2021

Abstract

We established an evaluation index of the rock mass stress state for underground coal mines using the strength-stress ratio based on the measured in situ stress and the generalized Hoek–Brown strength criterion. Three in situ stress states, σ_cm/σ_1m < 1.4 (high), 1.4 < σ_cm/σ_1m < 3.6 (medium), and σ_cm/σ_1m > 3.6 (low), were established based on the value of the unconfined compressive strength (σ_cm) and the maximum principal stress of the rock mass (σ_1m). This index classifies the Burtai mine as a medium-high in situ stress field, which is in agreement with the on-site situation, establishing the reliability of the index. The working face was a three-dimensional geological model based on the log sheets. The initial conditions for the model were determined using the combined measurements of the in situ stress regression model. We performed numerical simulations of the roof stress field distribution under repeated mining. Mining the overlying coal seam leads to significant variation in the value and direction of the main roof, σ₁, within the range of the front abutment pressure under the pillar and gob. Along the main roof strike direction, σ₁ under the pillar is 1.5 times that under the gob, and the σ₁ direction under the pillar is deflected by 5°, which is 30° smaller than that under the gob. This provides a reference for optimized underground coal mining.

1. Introduction

The combination of in situ stress and mining stress in underground coal mines forms a mining stress field. This stress field is directly related to the stress state of the overlying strata of the coal seam and the structural movement characteristics, and it affects the safe mining of the working face. The cause of in situ stress is complex, and it has an irregular distribution. It is affected by several factors, including depth, topography, structure, lithology, and residual stress. Hast [1] studied the distribution of in situ stress in different regions of western European countries, such as Norway and Sweden, based on which Stephansson et al. [2] established a regional in situ stress database in 1969. Brown and Hoek [3] analyzed the results of in situ stress in the shallow crust at a depth of 3 km in 1978. Haimson [4] summarized the in situ stress distribution in the United States at a depth of 0–5 km in 1980. Sheorey [5] proposed a crustal stress calculation model suitable for 138.87 m∼33.73 km. The World Stress Map (WSM) project [6] compiles the direction of the maximum horizontal principal stress (σ_H). WSM 2016 has 42,870 data records, which is double that of WSM 2008. Statistical studies on the distribution of crustal stress fields have been conducted in mainland China [7–12]. Kang et al. [13] compiled 1,357 in situ stress data in the “Database of in situ stress in underground coal mines in China,” and observed three stress field distributions: σ_H > σ_h > (<150 m), σ_H > > σ_h (150–1000 m), and > σ_H > σ_h (>1000 m). Yang [14] investigated the in situ stress in 26 groups of the northern Ordos Basin, China, and analyzed the principal stress distribution with depth, σ_H = 5.3367 ± 0.023 H, correlation coefficient R² = 0.94, and σ_h = 0.6559 ± 0.017 H, R² = 0.55; the dominant orientation of σ_H is the NEE. However, it cannot be used for establishing mine-scale engineering practice guidelines due to the variation in survey locations, and therefore, area-specific studies on in situ stress measurement and application are required.

The in situ stress state and the strength of the intact rock mass are essential for the stability of underground coal mines. At present, the empirical method, the strength-stress ratio method, the stress ratio method, and the geological marker method are used for evaluating the in situ stress state. Although 20 MPa is the typical high-stress standard, Yu et al. [15] proposed a standard of 0–10 MPa (low stress), 10–18 MPa (medium stress), 18–30 MPa (high stress), and greater than 30 MPa (superhigh stress). The strength-stress ratio method is the most widely used method [16, 17], but it lacks a unified standard. The Sheorey model [5] has been widely recognized in engineering design since it can predict horizontal stress and address the discrete stress values. The evaluation parameters include weathered rock mass or shallow surface engineering areas, which do not conform to the actual condition of coal mines [18–22]; therefore, the stress state evaluation indicators based on in situ stress data are required.

The strength-stress ratio method was used to establish the evaluation index for the stress state of the mine rock mass based on the in situ stress measurement through Kaiser effect and the generalized Hoek–Brown rock mass strength criterion, and a preliminary engineering verification was carried out. A refined geological model for the stratum occurrence conditions in the Burtai mine was established with 42# coal seam mining combined with the measured in situ stress regression model. The roof stress field of the working face under repeated mining conditions provided a reference for optimizing the mining plan in the Burtai mine and provided the guidelines for mines under similar conditions.

2. In Situ Stress Measurement in the Study Area

The Burtai mine is an extra-large mine in the Shendong mining area in China with a designed production capacity of 20 Mt/a and a service life of 71.9 a. Currently, 22# coal seam, having a thickness between 0.82 and 5.80 m and an average thickness of 2.52 m, and 42# coal seam, with thickness between 0.90 and 7.68 m and 6.50 m average thickness of the coal seam, are mined in the entire area. The coal seam has a simple structure with a stable layer, and it is a low-gas mine. The upper and lower coal seams of 22# and 42# are mined at the same time. With the increase in mine depth beyond 400 m, in situ stress becomes prominent due to increased mining intensity. Repeated mining causes severe disturbance with complex breaking characteristics of the overlying rock. For example, roof sinks and the bottom heaves were observed in the return roadway along the 22103 working face, resulting in the roadway being re-dug twice. The slab was damaged in the 0.2–1.5 m depth along the 42105 working face because of mechanized mining and poor caving technology, and roof weighting over a large area was observed. The above phenomena are inseparable from the in situ stress field.

Yang [14, 23, 24] established an in situ stress test based on the Kaiser effect of a borehole core and analyzed the in situ stress distribution in and around the boreholes of the Burtai mine, as shown in Figure 1.

(a)

(b)

The principal stress values are discrete but increase with an increase in depth (Figure 1(a)) and the fitting equations are as follows (equation (1)). Three stress field distributions are present: σ_H > σ_h > (<172 m), σ_H > > σ_h (172–800 m), and > σ_H > σ_h (>800 m).

In Figure 1(b), the variation of k ((σ_H + σ_h)/2) with depth is shown in equation (2) and compared with the k envelope (equation (3)) obtained by Brown and Hoek. The variation is similar to the progressive value of the inner envelope, and the k range reduces with an increase in depth (at 350 m, k = 0.98).

3. Evaluation of the Stress State in Underground Coal Mining

3.1. Evaluation Index of the Stress State

The rock mass stress state affects the rock mass stability and includes in situ stress, rock mass strength, and rock mass structural quality. Several classification systems have been developed for rock mass. The rock quality designation (RQD), proposed by Deere in 1964 [25], is one of the simplest systems of rock mass classification. The rock mass rating system (RMR) [26] and tunneling quality index (Q system) [27] are the most widely utilized rock mass classification schemes, and they have been widely adopted in the mining industry. The Hoek–Brown strength criterion, developed in 1980 [28–31], can account for the nonlinear behavior of the ultimate principal stresses when a rock mass fails and describes the influence of the rock mass structure and stress state on its strength. The parameters can be obtained qualitatively or quantitatively using conventional mechanical tests and field structure descriptions. The associated geological strength index (GSI) is widely used for estimating the strength and deformation of heavily jointed rock masses. The stress-based failure criteria developed by Coulomb, Mohr, and Hoek and Brown are extensively applied in rock mechanics. The Hoek–Brown strength criterion has undergone several modifications, and the final version (2018) is given by the following equation and described by Hoek and Brown in 2019:where m_b, s, and α are the rock mass material constants, given bywhere σ₁ and σ₃ are the maximum and minimum principal stresses at failure, respectively; σ_ci is the unconfined compressive strength; m_b is a rock property, modified by the GSI and degree of disturbance of the rock mass, respectively; s is a constant determined by the GSI and degree of disturbance of the rock mass; α is a constant determined by the GSI; m_i, s = 1, and α = 0.5 are the material constants for intact rock; and D is a factor that depends upon the degree of disturbance to which the rock mass has been subjected to blast damage and stress relaxation.

When σ_tm < σ₃ < σ_ci/4, the expression is given by the following equation:

Incorporating equation (6) into the Hoek–Brown strength criterion, let , σ_ci = k_mσ_cm. Then,

The relationship between the σ_cm/σ₁ ratio and GSI for different friction coefficients is shown in Figure 2. With an increase in the friction coefficient, the σ_cm/σ₁ ratio changes significantly. When 0.2 < μ < 0.6, σ_cm/σ₁ ratio is between 0.01 and 0.3, when 0.6 < μ < 1.0, σ_cm/σ₁ ratio lies between 0.05 and 0.7, and the σ_cm/σ₁ ratio generally varies between 0.05 and 0.5.

(a)

(b)

(c)

(d)

(e)

(f)

The relationship between the σ_cm/σ₁ ratio and GSI for different disturbance coefficients is shown in Figure 3. With a decrease in GSI, the variation of the disturbance coefficient (D) increases, indicating that the deteriorating geological conditions increase the degree of disturbance in the in situ rock masses because of blasting, rock mass excavation, and rock mass unloading.

(a)

(b)

The in situ deformation tests of 50 groups of rock masses indicate that D around the prospecting drift is 0.25 [32], while D caused by the excavation of the mining roadway is less than 0.4. When μ = 0.3 and m_i = 24, the ratio of roadway excavation to σ_cm/σ₁ does not exceed 0.02; when μ = 0.6 and m_i = 24, the ratio of roadway excavation to σ_cm/σ₁ exceeds 0.1. Therefore, the mining roadway is less sensitive to D compared to GSI.

The coal-measure strata are clastic sedimentary rocks, mainly sandstone. The values of m_i for various rocks are as follows: siltstone 7 ± 2, hard sandstone 18 ± 3, sandstone 17 ± 4, conglomerate 21 ± 3, and breccia 19 ± 5 [30, 31]. Most mined sandstones have a core length greater than 10 cm, and the joints are weakly weathered or non-weathered. The GSI of the surrounding rock of the coal seam is above 30–40. The friction coefficient of sandstones ranges between 0.1 and 0.6, with an average of 0.3. The σ_cm/σ₁ ratio of sandstone in most coal-measure strata lies between 0.1 and 0.4 for different friction coefficients. The higher boundary value of the σ_cm/σ₁ ratio is suitable for most rock mass stress state evaluation of coal-measure strata. The rock mass stress will significantly impact its stability for σ_cm/σ₁ < 0.4 and may cause rock mass instability, which is defined as a state of rock mass with high stress.

Stress concentrations occur in the rock mass around the roadway during underground mining. Therefore, the stress concentration factor should be considered for calculating the maximum principal stress in the rock mass; hence, σ₁ can be expressed aswhere η is the maximum stress concentration factor of the rock mass around the roadway and σ_1m is the measured maximum principal stress.

The rock surrounding the roadway is a nonhomogeneous, discontinuous, and nonlinear material with complicated loading and boundary conditions. For a circular roadway, the maximum stress concentration factor is 2.0 and 3.0 under uniform and nonuniform stress field conditions, respectively. The maximum stress concentration factor for a rectangular roadway stays below 4.0 and is generally less than 3.5. Under most coal seam and mining conditions, the front abutment pressure of the working face is in the range of 2.5–3.5 [33]. At η = 3.5, the rock mass starts to fail for σ_cm/σ_1m < 1.4, which corresponds to a high in situ stress state. The corners where the two working faces meet experience stress concentration seven times the vertical stress of the original rock, or even higher. The rock mass failure occurs locally in the roadway section for η = 9.0 and 1.4 < σ_cm/σ_1m < 3.6, and the stress state corresponds to the middle in situ stress state. When σ_cm/σ_1m > 3.6, local rock mass collapses, or the structural plane becomes unstable because it is difficult to achieve a ninefold stress concentration factor in coal-measure strata, and this corresponds to a low in situ stress state. Based on the strength-stress ratio, the evaluation index for the rock mass stress state of coal mines is presented in Table 1.

σ_cm is the unconfined compressive strength of the original rock mass of the surrounding rock of the roadway and can be estimated using the Hoek–Brown strength criterion (equation (4)) or using RocData software; σ_1m is the maximum principal stress in the in situ stress field of the surrounding rock of the roadway, which can be calculated using the theoretical model of in situ stress, or obtained using in situ stress testing.

3.2. Engineering Verification

The relationship between σ_cm/σ_1m and the depth of all rock formations in the Burtai mine is used to verify the in situ stress evaluation index of the rock mass strength-stress ratio (Table 1) as shown in Figure 4(a). The roof and floor data around the 22#, 42#, and 52# down coal seams are shown in Figure 4(b), where σ_ci is the average uniaxial compressive strength of the rock, σ_cm is the average rock mass strength, and σ_1m is the measured maximum principal stress. GSI is determined using a combination of the rock quality designation (RQD), the rock block index (RBI), and the core recovery percentage (CRP), with the disturbance coefficient D set to 0. The parameters are listed in Tables 2 and 3.

(a)

(b)

Figure 4(a) shows 89 measuring points in the Burtai mine, of which 59, with σ_cm/σ_1m < 1.4, account for 66.3% of the total; 27, with 1.4 < σ_cm/σ_1m < 3.6, account for 30.3%; and 3, with σ_cm/σ_1m > 3.6, account for 3.4%. In Figure 4(b), among the 24 measuring points in Burtai mine, 11, with σ_cm/σ_1m < 1.4, account for 45.8% of the total; 12, with 1.4 < σ_cm/σ_1m < 3.6, account for 50.0%; and 1, with σ_cm/σ_1m > 3.6, accounts for 4.2%. The average values of σ_cm/σ_1m of the roof and floor around 22#, 42#, and 52# down coal seams were 1.59, 1.25, and 2.00, respectively. The groups with σ_cm/σ_1m < 1.4 are divided according to the depth into 350–450 m, 450–550 m, and 550–650 m, which account for 37.5%, 50.0%, and 42.9% of the total, respectively. The ratio of σ_cm/σ_1m < 1.4 increases with an increase in the depth, indicating the Burtai mine to be a medium-high in situ stress field.

According to the empirical method, 39 groups with σ_h < 10 MPa account for 43.8% of the total; 42 with 10 < σ_H < 18 MPa account for 47.2%; and 8 with σ_H > 18 MPa account for 9%. Among them, 34 groups were above 350 m, of which 29 groups with σ_h < 10 MPa, account for 85.3%; 55 groups were below 350 m, of which 11 groups with σ_h < 10 MPa account for 20%; and 65.5% groups have 10 < σ_h < 18 MPa. Therefore, the in situ stress field in the Burtai mine is low in situ stress field above 350 m and medium in situ stress field below 350 m. According to the stress ratio method, there are 86 groups with (σ₁ + σ₂ + σ₃)/σ_V > 2, accounting for 96.6% of the total, indicating the majority of the Burtai mine as high in situ stress field, which does not conform to the actual condition of the Burtai mine.

The current mining depth of 42# coal seam in the Burtai mine is greater than 400 m, resulting in nonlinear dynamic phenomena such as bottom heave, slab, and large-area roof weighting. With a further increase in mining depth, the roof and floor will be under high in situ stress along with an increase in the accumulated elastic energy. This justifies the medium-high in situ stress field of the Burtai mine and necessitates the active adoption of the mine pressure prevention technology, and establishes the reliability and applicability of the strength-stress ratio index.

4. Numerical Simulation of the Stress Field in Repeated Mining of Coal Seams

4.1. Model Construction

We analyzed six log sheets (E014, E019, E026, E030, E051, and E060) around the 42107 working face in the Burtai mine corresponding to the SJ–3 log sheet on the 42105 working face [34]. The representative marker layer, which roughly divides the Burtai mine into 15 strata from top to bottom, was selected as the strata interface. A refined three-dimensional geological model of the working face was constructed using Rhinoceros software and imported into Flac3D using griddle software (Figure 5). The spatial position of the working face is illustrated in Figure 6. The numerical model of the working face was length × width × height = 800 m × 900 m × 600 m, and it was divided into 1,605,551 units and 283,006 nodes. The initial conditions of the model are determined in combination with the measured in situ stress regression model (equation (1)).

The physical and mechanical parameters of the model strata were determined using laboratory tests of the coal and rock in the SJ–3 borehole at 42105 working face in the Burtai mine [34]. The tensile strength of rock mass (σ_tm) is calculated by RocLab 1.0 software (Table 4).

The rock follows the Mohr–Coulomb elastoplastic model, while the broken rock mass in the caving zone follows the double-yield-surface elastoplastic model. Based on the compaction characteristics of the broken rock mass, the stress-strain relationship of the caving zone rock mass aligns well with the actual conditions of the coal mine and is given by the following equation:where is the vertical stress of the caving zone in the gob (MPa); ε is the vertical strain, ε_m is the maximum vertical strain, taken as ε_m = (b – 1)/b; b is the comprehensive crushing expansion coefficient of the rock mass in the caving zone, calculated using the caving zone H_c = h/(b – 1); E₀ is the initial elasticity modulus of the rock mass, MPa, calculated using E₀ = 10.39σ_c^1.042/b^7.7 [35]; and σ_c is the uniaxial compressive strength (UCS) of the rock mass in the caving zone, MPa.

The UCS of the roof above 22# and 42# coal seams are greater than 40 MPa, indicating them to be hard rock strata. According to the coal mining conditions, the average thickness of coal seam 22# is 3.6 m, and the average thickness of the 42# coal seam, which uses mechanized mining with sublevel caving technology, is 6.5 m. The machine mining height is 3.6 m, the coal caving height is 2.9 m, and the top coal caving rate is 80%, indicating the actual mining height of the 42# coal seam to be 5.92 m. The heights of the caving zone (H_c) and the fissure zone (H_f) can be calculated using equations (10) and (11) and Table 5 [35–37]. The heights of the two zones formed after the 22# and 42# coal seam working faces with longwall mining are 15.3 m, 57.0 m, and 20.8 m; and 65.0 m, 1.236 and 1.284, respectively; and ε_m is 0.19 and 0.22, respectively. The mechanical parameters of the caving zone rock mass in the gob are tabulated in Table 6.

4.2. Results and Analysis

The impact of the overlying pillar on the mining face was determined using simulations of simultaneous upper and lower coal seam mining. The widths of overlying 22# coal seam pillars are 5 m, 15 m, 25 m, 35 m, and 45 m. Coal seam 42# lags the 22# coal seam by 300 m, and the excavation distance is 300 m. Two survey lines (117 m and 200 m from the return roadway) on strike and one survey line (5 m from the rib) in an inclined direction are arranged on the main roof of the 42107 working face at the height of 25 m from the working face. The principal stress distribution of the main roof under different pillars and the gob is monitored, as shown in Figure 7.

(a)

(b)

(c)

(d)

4.2.1. Maximum Principal Stress Distribution on the Strike

σ₁ of the main roof (380 m deep) above 42# coal seam increases by 10 MPa when advancing 350 m because of the overlying pillars of 22# coal seam (Figures 7(a) and 7(c)). The principal stress increases by different degrees with the increase in the width of the pillar from 300 m. The larger the pillar width, the faster the increase in stress. The increase rate increases from 60 m advances and reaches the maximum for 4–9 m advances. The rate decreases rapidly above and behind the working face and reduces to 6 MPa and 30 m behind the gob. Original rock stress zones, stress increase zones, and stress reduction zones are present on the strike of the working face under the pillar, and the size of the stress increase zone is different under different pillars. Although σ₁ increases with an increase in pillar width, the rate of increase reduces, while the principal stress influence ranges remain unchanged. The peak value distance from the rib is approximately 3 m, and the stress peak coefficients are 1.95, 2.19, 2.31, 2.35, and 2.45, respectively, under different pillar widths. The distribution of σ₁ on strike under the gob is similar to that under the pillar, and the peak coefficients are reduced to 1.32, 1.41, 1.47, 1.56, and 1.66, respectively.

The roof σ₁ direction of the working face under the pillar is first deflected towards the z-axis in the advanced range and then deflected in the reverse direction (Figures 7(b) and 7(d)). The angle between the σ₁ direction and the z-axis is 6° for advancing 350 m, and it is biased towards the gob. The angle increases to 13° from advancing 300 m and gradually decreases to 3° for 50 m. The angle remains constant with the increase in pillar width, and the reverse deflection occurs within advancing 50 m, reaching 30° above the working face, and 55° at 30 m behind the gob. The deviation of the σ₁ direction below the pillar remains constant beyond 350 m and is subsequently affected by the overlying coal mining. The angle between the σ₁ direction and the z-axis increases sharply to 68° within advancing 350 m to 250 m. Subsequently, reverse deflection occurs, and the angle decreases. The deflection angle ranges between 50° and 80° and decreases with an increase in pillar width, with a sharp decrease to 35° within advancing 16 m. After the working face, the σ₁ direction continues to deflect to the gob, and the deflection reaches 60° for 25 m behind the gob. The main roof σ₁ deflection angle within 25 m in front of the rib under the gob is 35° because of the overlying coal mining and is 10° larger than solid coal. The friction coefficient of rock is distributed between 0.1° and 0.8°, and the predominant fracture angle is 25–45°, as seen from the Griffith physical model of closed microcrack propagation under pressure [38]. The shear-type (II) and tension-shear mixed-type (I-II) expansion of the coal and rock mass cracks are observed along the vertical plane within the severely affected range of mining. The principal stress of the roof remains unchanged, and the principal stress direction is rotated according to the conditions of the tensile type (I) expansion of the roof cracks [38]. When σ₁ rotates to the dominant angle, the maximum tensile stress reaches the tensile strength of the roof, and the microcracks in the top coal expand. The expansion of the roof (top coal) cracks can be attributed to the principal stress rotation, the increase in σ₁, and the decrease in σ₃.

4.2.2. Maximum Principal Stress and Deviator Stress in the Inclined Direction

The principal stress distribution of the main roof has a “single-peak” shape (Figure 8) along the inclined direction and 5 m in front of the rib with high values under the pillar and low values on both sides. With an increase in the pillar width, the peak coefficients of σ₁ increase, the range of increase reduces, and the severe influence range increases. The principal stress peak coefficients are 1.91, 2.11, 2.26, 2.35, and 2.43, respectively, under different pillar widths of 5 m, 15 m, 25 m, 35 m, and 45 m, and the severe influence range is between 100 and 150 m. Both sides of the working face are affected by the overlying pillar. Although the principal stress increases with an increase in the pillar width, the increase is much smaller than that under the pillar. The σ₁ of the main roof on the strike survey line under the pillar is 1.5 times that under the gob. Similar to the σ₁ distribution, the deviator stress on inclined direction also presents a “single-peak” shape, with peak values of 15.8, 18.1, 19.8, 20.7, and 21.4 MPa, respectively, under different pillar widths of 5 m, 15 m, 25 m, 35 m, and 45 m.

(a)

(b)

The deviator stress is the largest within the severe influence range of the pillar and fluctuates significantly. The main roof in the front of the rib breaks under tensile stress because of considerable accumulated strain energy, which transforms into the fracture surface energy and the initial velocity of the fractured rock block, resulting in an impact load. Roof fracture and instability can cause a dynamic catastrophe of the working face due to the combined dynamic and static load.

5. Conclusion

The strength-stress ratio method was used to establish the evaluation index for the stress state of the mine rock mass based on the in situ stress measurement through Kaiser effect and the generalized Hoek–Brown rock mass strength criterion, and a preliminary engineering verification was carried out. A refined geological model for the stratum occurrence conditions in the Burtai mine was established with 42# coal seam mining combined with the measured in situ stress regression model. The roof stress field of the working face under repeated mining conditions is studied, and the main conclusions are as follows:(1)An evaluation index of the rock mass stress state for an underground coal mine was established using the strength-stress ratio. It corresponds to high in situ stress state for σ_cm/σ_1m < 1.4, medium in situ stress state for 1.4 < σ_cm/σ_1m < 3.6, and low in situ stress state for σ_cm/σ_1m > 3.6.(2)The index is applied in the Burtai mine and classifies it as a medium-high in situ stress field, which is in agreement with the on-site situation, establishing the reliability of the index.(3)The value and direction of the main roof σ₁ within the range of the front abutment pressure under the pillar and gob vary significantly because of the mining of the overlying coal seam. The value of σ₁ under the pillar along the main roof strike direction is 1.5 times that under the gob, and the σ₁ direction under the pillar is deflected by 5°, which is 30° smaller than that under the gob. The rotation of principal stress direction, the increase in σ₁, and the decrease in σ₃ can cause the expansion of the roof (top coal) cracks.

Data Availability

The datasets used during the study are available from the corresponding author upon request.

Conflicts of Interest

The authors declare no conflicts of interest.

Acknowledgments

This study was supported by the Fundamental Research Program of Shanxi Province (202101D221355), Project 1331 of Shanxi Province, Graduate Education Innovation Program of Shanxi Province and Shanxi Datong University (2021YJJG306 and 21JG13), the Key Research Development Program of Shanxi Province (201903D121070), the Key Research Development Program of Datong City, Shanxi Province (2019013), Datong Basic Research Project (2019159), and Shanxi Datong University Research Project (2020Q017) and Doctoral Research Startup Fund.

References

N. Hast, “The state of stress in the upper part of the earth’s crust,” Tectonophysics, vol. 8, no. 3, pp. 169–211, 1969.
View at: Publisher Site | Google Scholar
O. Stephansson, P. Särkkä, and A. Myrvang, “State of stress in Fennoscandia,” in Proceedings of the Isrm International Symposium, Cappadocia, Turkey, August 2016.
View at: Google Scholar
E. T. Brown and E. Hoek, “Trends in relationships between measured in-situ stresses and depth,” International Journal of Rock Mechanics and Mining Science & Geomechanics Abstracts, vol. 15, no. 4, pp. 211–215, 1978.
View at: Publisher Site | Google Scholar
B. C. Haimson, “Near-surface and deep hydrofracturing stress measurements in the Waterloo quartzite,” International Journal of Rock Mechanics and Mining Science & Geomechanics Abstracts, vol. 17, no. 2, pp. 81–88, 1980.
View at: Publisher Site | Google Scholar
P. R. Sheorey, “A theory for in Situ stresses in isotropic and transverseley isotropic rock,” International Journal of Rock Mechanics and Mining Science & Geomechanics Abstracts, vol. 31, no. 1, pp. 23–34, 1994.
View at: Publisher Site | Google Scholar
M. L. Zoback, “First‐and second‐order patterns of stress in the lithosphere: the World Stress Map Project,” Journal of Geophysical Research Solid Earth, vol. 97, no. B8, pp. 11,703–11,728, 1992.
View at: Google Scholar
S. X. Yang, Study on the Distribution Characteristics of Crustal Stress Field in Chinese Mainland, Beijing Jiaotong University, Beijing, China, 2013.
F. Jing, Research on the Distribution Rule of the Shallow Crustal Geostress Field in the China Mainland and Engineering Disturbance Characteristics, Graduate School of Chinese Academy of Sciences (Wuhan Institute of Geotechnical Mechanics), Wuhan, China, 2009.
C. J. Feng, Study on the Present in Situ Stress Field beneath the Capital Circle Region, Chinese Academy of Geological Sciences, Beijing, China, 2014.
H. P. Kang, J. Ling, and L. X. Yan, “Study on distribution characteristics of underground in–situ stress field in Shanxi coal mining area,” Chinese Journal of Geophysics, vol. 52, no. 7, pp. 1782–1792, 2009.
View at: Google Scholar
H. C. Zhu and Z. Y. Tao, “In situ stress distribution in different rocks,” Acta Seismologica Sinica, vol. 16, no. 1, pp. 49–63, 1994.
View at: Google Scholar
D. A. Zhao, Z. M. Chen, and X. L. Cai, “Statistical analysis of distribution law of in–situ stress field in China,” Chinese,” Journal of Rock Mechanics and Engineering, vol. 26, no. 6, pp. 1265–1271, 2007.
View at: Google Scholar
H. P. Kang, B. D. Yi, and F. Q. Gao, “Underground geostress database and geostress distribution law of coal mines in China,” Journal of China Coal Society, vol. 44, no. 1, pp. 23–33, 2019.
View at: Google Scholar
D. H. Yang, Research on In–Situ Stress Measurement Method and Application Based on Kaiser Effect of Borehole Core, China University of mining and Technology, Beijing, China, 2019.
X. F. Yu, Y. R. Deng, and H. H. Liu, Stability Analysis of Surrounding Rock in Underground Engineering, Coal Industry Press, Beijing, China, 1983.
L. S. Xu, B. M. Tang, and C. C. Mu, “Research status of problems related to high ground stress and rockburst,” Technology of Highway and Transport, vol. 4, no. 12, pp. 48–51, 2002.
View at: Google Scholar
C. H. Wang, Q. L. Guo, and L. Jia, “Theoretical analysis of high stress criterion based on Hoek Brown strength criterion,” Progress in Earthquake Sciences, vol. 32, no. 6, pp. 3325–3332, 2012.
View at: Google Scholar
D. Yang, Y. Zhao, Z. Ning, Z. Lv, and H. Luo, “Application and development of an environmentally friendly blast hole plug for underground coal mines,” Shock and Vibration, vol. 2018, Article ID 6964386, pp. 1–12, 2018.
View at: Publisher Site | Google Scholar
Q. Wu, X. Li, L. Weng, Q. Li, Y. Zhu, and R. Luo, “Experimental investigation of the dynamic response of prestressed rockbolt by using an SHPB-based rockbolt test system,” Tunnelling and Underground Space Technology, vol. 93, Article ID 103088, 2019.
View at: Publisher Site | Google Scholar
Q. Wu, L. Chen, B. Shen, B. Dlamini, S. Li, and Y. Zhu, “Experimental investigation on rockbolt performance under the tension load,” Rock Mechanics and Rock Engineering, vol. 52, no. 11, pp. 4605–4618, 2019.
View at: Publisher Site | Google Scholar
T. Zhou, J. Zhu, and H. Xie, “Mechanical and volumetric fracturing behaviour of three-dimensional printing rock-like samples under dynamic loading,” Rock Mechanics and Rock Engineering, vol. 53, no. 6, pp. 2855–2864, 2020.
View at: Publisher Site | Google Scholar
T. Zhou, J. B. Zhu, Y. Ju, and H. P. Xie, “Volumetric fracturing behavior of 3D printed artificial rocks containing single and double 3D internal flaws under static uniaxial compression,” Engineering Fracture Mechanics, vol. 205, pp. 190–204, 2019.
View at: Publisher Site | Google Scholar
D. H. Yang, Y. X. Zhao, and C. Zhang, “Experimental study on the influence of cyclic loading on Kaiser effect of sedimentary rocks,” Chinese Journal of Rock Mechanics and Engineering, vol. 37, no. 12, pp. 2697–2708, 2018.
View at: Google Scholar
D. H. Yang, Z. X. Ning, and Y. M. Li, “In situ stress measurement and analysis of the stress accumulation levels in coal mines in the northern Ordos Basin, China,” International Journal of Coal Science & Technology, vol. 4, 2021.
View at: Google Scholar
D. U. Deere, “Technical description of rock cores for engineering purposes,” Rock Mechanics and Engineering Geology, vol. 1, no. 1, pp. 16–22, 1964.
View at: Google Scholar
Z. T. Bieniawski, “Geomechanics classification of rock masses and its application in tunnelling,” in Proceedings of the 3rd Congress of International Society Rock Mechanics, Denver, CO, USA, September 1974.
View at: Google Scholar
N. Barton, R. Lien, and J. Lunde, “Engineering classification of rock masses for the design of tunnel support,” Rock Mechanics Felsmechanik Mcanique des Roches, vol. 6, no. 4, pp. 189–236, 1974.
View at: Publisher Site | Google Scholar
E. Hoek and E. T. Brown, Underground Excavations in Rock, Institution Mining and Metallurgy, London, UK, 1980.
E. Hoek, C. Carranza–Torres, and B. Corkum, “Hoek–Brown failure criterion–2002 edition,” in Proceedings of the Fifth North American Rock Mechanics Symposium, Toronto, Canada, July 2002.
View at: Google Scholar
E. Hoek and E. T. Brown, “Practical estimates of rock mass strength,” International Journal of Rock Mechanics and Mining Sciences, vol. 34, no. 8, pp. 1165–1186, 1997.
View at: Publisher Site | Google Scholar
E. Hoek and E. T. Brown, “The Hoek–Brown failure criterion and GSI–2018 edition,” Joumal of Rock Mechanics and Geotechnical Engineering, vol. 11, no. 3, pp. 445–463, 2019.
View at: Google Scholar
Y. H. Song and L. Cong, “Estimation of deformation modulus and disturbance coefficient of rock mass based on GSI,” Journal of Basic Science and Engineering, vol. 22, no. 1, pp. 27–34, 2014.
View at: Google Scholar
M. G. Qian, P. W. Shi, and J. L. Xu, Mining Pressure and Strata Control, China University of mining and Technology Press, Xuzhou, China, 2010.
H. M. Li, H. G. Li, and S. S. Peng, Physical and Mechanical Properties of Rocks in Shendong Mining Area, Science Press, Beijing, China, 2018.
M. Salamon, “Mechanism of caving in longwall coal mining,” in Proceedings of the 31st US Symposium of Rock Mechanics, pp. 161–168, Golden, CO, USA, June 1990.
View at: Google Scholar
L. S. Jiang, Q. S. Wu, and X. Y. Li, “Study on coupling analysis method of mining stress and compaction bearing capacity of goaf,” Journal of China Coal Society, vol. 8, pp. 1951–1959, 2017.
View at: Google Scholar
H. Yavuz, “An estimation method for cover pressure re-establishment distance and pressure distribution in the goaf of longwall coal mines,” International Journal of Rock Mechanics and Mining Sciences, vol. 41, no. 2, pp. 193–205, 2004.
View at: Publisher Site | Google Scholar
A. A. Griffith, “The phenomena of rupture and flow in solids,” Philosophical Transactions of the Royal Society of London, vol. A221, pp. 163–198, 1921.
View at: Google Scholar

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Copyright © 2021 Donghui Yang et al. This is an open access article distributed under the Creative Commons Attribution License, which permits unrestricted use, distribution, and reproduction in any medium, provided the original work is properly cited.

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