Case Study on the Mechanism of Influence of Stoppage on Ground Pressure under Different Rates of Advance

Gao, Xiaoxu; Shi, Xinyu; Guo, Weibin

doi:https://doi.org/10.1155/2021/5574693

Advances in Civil Engineering

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Development of Ground Damage and its Correlation with Underground Mining

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Volume 2021 | Article ID 5574693 | https://doi.org/10.1155/2021/5574693

Case Study on the Mechanism of Influence of Stoppage on Ground Pressure under Different Rates of Advance

Xiaoxu Gao,^1,2Xinyu Shi ,^1,3and Weibin Guo^1,2

Academic Editor: Dezhong Kong

Received28 Feb 2021

Accepted13 Jul 2021

Published22 Jul 2021

Abstract

Because of daily maintenance, equipment damage, gas overrun, and other force majeure factors, the continuous stopping of the working face causes the roof pressure to accumulate, which leads to causing accidents such as coal wall spalling and roof falling. To address the roof safety problem caused by continuous stoppages, the 620 working face in the Huangling mining area is taken as the research object. Through field measurement, theoretical analysis, numerical simulation, and other research methods, the influence and mechanism of stopping pressure under different rates of advance are studied. The results show that the velocity factor of roof load transfer is positively correlated with the advancing velocity of the working face; the reasonable length of the suspended roof is mainly affected by the number of caving holes and the effect of pressure relief; and comparing the two stages of advance speed of 4.8 m/d and 12.8 m/d, the periodic weighting step distance of the latter increases by 24.4% compared with the former, and the rate of increase of support load caused by stopping mining increases by 42.1% compared with the former. The roof pressure accumulation caused by stopping mining is increased. Taking appropriate measures for local forced caving of the working face can release the roof pressure and reduce the risk of local caving of the working face. The study can provide a theoretical basis for roof control of continuous stopping under similar engineering conditions.

1. Introduction

Due to force majeure factors such as daily maintenance of working face, equipment damage, special geological structures, and gas overrun, the continuous stoppage of mining in the normal advancing process of working face is inevitable, and the characteristics of ground pressure show obvious differences under different rates of advance [1]. In this regard, Xie [2–4] analysed the failure field and stress field of the fully mechanised top coal caving face under different rates of advance through numerical simulation and similar simulation methods. The results show that when the unit mining depth increases, the extent of failure zones in the rock around the working face decreases, but a large amount of energy accumulates inside the rock mass, and the possibility of local rock burst increases; according to Yang and Liu [5, 6], the integrity of surrounding rock and the volume of broken rock block are positively correlated with the rate of advance of working face in shallow coal seams; S. Yang and J. Yang [7–13] believed that, in high-intensity mining, the rate of occurrence of coal and rock disasters is affected by the rate of advance and the working face length. The greater the coal rock disaster rate, the greater the first weighting step of the working face roof; Zhu and Xu [14, 15] analysed the damage to a coal and rock mass under different geological conditions by combining numerical simulation and field measurement methods. The analysis results show that the faster the working face advances, the greater the stress concentration on both sides of the working face, and the extent of the plastic zone is reduced. These research results used weighting characteristics of working face and coal and rock catastrophes under different rates of advance but fail to reveal the ground pressure on the working face when the mining is stopped. Due to special reasons, if the advance distance has not reached the collapse step distance, a large amount of energy accumulated in the roof cannot be released, the risk of a roof fall in the working face is greatly increased, and, at the same time, the hydraulic support movable column shrinks and the hydraulic pipe bursts. In view of this kind of engineering problem, the author, by means of numerical simulation and field measurement, took the 620 working face of a mine in the Huangling mining area as the research background and investigated the ground pressure behaviour upon continuous stoppage at the working face in the process of high-speed advance thereof.

At present, the 620 working face of No. 2 Coal Seam in Huangling coal mine is being mined. The average thickness of the coal seam is 2.4 m, the dip angle is 0° to 6°, the average burial depth of the working face is 380 m, the width is 235 m, the advancing length is 2267 m, the current daily rate of advance is 12.8 m/d, the working face relies on the ZY6800/11.5/24d hydraulic support to automatically support the roof, the rated support resistance is 6800 kN, and the support is equipped with a PM32 electrohydraulic control system, which can record the live column load in real time. The roof and floor of the coal seam are thus as follows: the main roof is siltstone and fine sandstone and the thickness is 11.8 m; the direct roof is fine sandstone and the thickness is 8.7 m; and the direct bottom is mudstone and the thickness is 2.8 m. Due to force majeure factors such as daily maintenance, equipment damage, and gas overrun, the phenomenon of continuous stoppage affects mine safety. For face 620, the phenomenon of support pressing and hydraulic pipe bursting occurs regularly during the stoppage of the 620 working face (Figure 1). The table of rock mechanics parameters is shown in Table 1.

(a)

(b)

2. The Influence of Working Face Ground Pressure under Different Rates of Advance

To analyse the influences of ground pressure and the performance characteristics of stoppage pressure under different rates of advance, two different advancing sections of the 620 working face are selected for in situ testing (Table 2 and Figure 2). The stoppage points marked in Figure 2 are caused by local gas overrun of the working face. According to an investigation of the mining area, when the rate of advance is 4.8 m/d and 12.8 m/d, the basis for the judgment of periodic weighting is as follows: the end resistance of circulation reaches 35 MPa and 25 MPa, respectively, the high-level gas drainage concentration increases instantaneously, and the coal wall spalling is severe. When the working face stops mining, the bearing degree of the support increases, but it has not reached the average load for periodic weighting. After stopping mining, the roof pressure will be released with the continuous advance of the working face.

(a)

(b)

Combined with the load conditions of support in Figures 2(a) and 2(b), the measured weighting characteristics are listed in Table 3. The measured results show that when the rate of advance is 4.8 m/d and 12.8 m/d, the periodic weighting step distance of the latter is 24.4% higher than that of the former, and the support load rise rate caused by stoppage of the latter is 42.1% higher than that of the former, and the risk of roof fall of working face is greatly increased. Therefore, avoiding the fluctuation of the mining speed of the working face can decelerate the accumulation of dangerous roof pressure on the working face and play a positive role in the control of the rock strata; by comparing the rise of the support load when stopping mining under the two rates of advance, it is found that when the rate of advance is large, the accumulated pressure on the roof is also high, and the risk of a local roof fall in the working face is relatively low. Appropriate measures should be taken to release the roof pressure to ensure safe production at the working face.

3. Theoretical Analysis

3.1. Speed Factor of Load Transfer

Some research results show that, with the increase of the rate of advance, the peak load increases and moves forward [16–18]; and when the rate of advance is rapid, the mining failure caused by the direct roof is incomplete, the roof load increase caused by mining cannot fully act on the support, and the support load remains low [19, 20]. On the other hand, when the main roof collapses, the main load is provided by the load layer on the top of the main roof. The load transfer process and the rate of advance have a time effect. That is, when the working face is advancing normally, the load transfer characteristics of the main roof are such that it unloads first, then arches, and after that unloads, until the original state of stress in the rock is restored. However, when the rate of advance is decelerated due to special reasons, secondary unloading will occur. For this reason, some scholars have proposed a time transfer factor KT [21–25], as given by the following:where P_Z is the load acting on the key block; K_r is the lithology factor; h₁ is the thickness of the loading layer; L is the length of the key block; and ρ represents the average bulk density of the key layer.

At the same time, the author thinks that the time factor K_t will change with time in the transformation process of the key blocks of the main roof. Therefore, the following formula can be obtained:where m is the time factor.

Based on the above results, a velocity transfer factor suitable for the main roof load in the Huangling mining area is proposed. The velocity transfer factor will change with the speed, and the following formula can be obtained:where n is the velocity factor.

During the periodic weighting of the main roof, when the rate of advance is kept constant, the periodic weighting step length is quasiconstant; therefore, it is considered that the periodic weighting step is a constant S_Z, and the load speed transfer factor at this time is deduced, as shown in the following formula:

When the rate of advance changes, the periodic weighting step distance of the main roof also changes. The greater the rate of advance, the larger the periodic weighting step distance; therefore, the correction coefficient m is introduced. According to the measured results of periodic weighting step distance of working face in the Huangling mining area, the value of correction coefficient m is set (it is positively correlated with the rate of advance of the working face). Taking the advance speed of the 620 working face as an example, m = 1.2, therefore, the load transfer speed factor applicable to the 620 working face is shown in the following formula:

Taking the load of the key block of the main roof, the extreme value of the speed factor for load transmission can be deduced as follows:

It can be seen from formula (6) that is jointly affected by periodic weighting step S_Z and lithologic factor K_r as determined by the thickness of load layer, lateral pressure coefficient, and internal friction angle, that is, given K_r, m, and n being constant, it can be considered that is positively correlated with periodic weighting step S_Z. At the same time, periodic weighting step S_Z is positively correlated with the rate of advance of the working face and load transfer factor is positively correlated with the rate of advance.

3.2. Reasonable Suspended Roof Length for Roof Pressure Relief

After the first fracture of the roof, with the continuous advance of the working face, one end of the roof rock beam is fixed on solid coal, and the other is suspended, showing a cantilever beam structure, and with the continuous advance of the working face, the roof periodically collapses [1, 26, 27]. If the roof load is evenly distributed, roof pressure will accumulate when stoppage occurs. Based on previous research results [28–30], the design relies on the drilling of several caving holes in the direct roof of the gateway to release the roof pressure and assist the roof caving. Two sections are made along the gateway and the working face, and the mechanical model of roof pressure relief fracture is drawn, as shown in Figure 3 [31–33].

(a)

(b)

It can be seen from Figure 3 that the load layer above the main roof exerts a uniform load Q on the main roof. If there are I caving holes under construction and the load released by each caving hole is Q, the support resistance provided by the support to the working face is the allowable support resistance [P]. According to the fact that the support can bear all the loads within the roof control area, thenwhere a is the distance of support control, L is the distance of hanging roof, and L = a + b.

Forced caving is used to ensure that the working resistance P provided by the support is not greater than the allowable support resistance [P] in case of periodic roof fracture. Then,

The reasonable length of the suspended roof based on the pressure relief of the caving hole and the allowable support resistance [P] of the support can be calculated:

According to formula (9), after the working face is determined, the roof control distance of the support and the allowable support resistance of the support are determined, and the main roof load is also determined. Therefore, the reasonable hanging roof length is mainly affected by the number of caving holes and the pressure relief effect.

4. Numerical Simulation Study

4.1. Establishment of Model

FLAC^3D finite difference simulation software is used here: the Mohr–Coulomb constitutive relationship is set for rock stratum and a strain-softening constitutive relationship is applied to model the coal seam. The mechanical parameters of rock mass are provided by the mine. To simplify the stratum, the strike direction is 345 m, the advancing direction length is 300 m, the vertical thickness is 81 m, and the thickness of the coal seam is 3 m. A total of 354,960 elements and 371,124 nodes are established. We limit the displacement around the model, fix the bottom, apply a 5 MPa uniform load to the top of the model, and assume a lateral pressure coefficient of 1.2. The numerical simulation model is shown in Figure 4. The numerical simulation mainly analyses the distribution of advance abutment pressure after the working face reaches a stable rate of advance. After the working face is advanced by 100 m, three measuring lines are established to monitor the abutment pressure distribution characteristics within 50 m ahead of the working face. The three monitoring lines are located on both sides of goaf and the middle of goaf, respectively. When simulating the advancing process of working face, the unit mining depth of the working face is replaced by time steps, and the difference of rate of advance is characterised by the difference in total operation steps [2, 7, 8]. Among them, the same excavation step distance and different excavation steps are used to replace the propulsion speed. The specific advancing scheme is summarised in Table 4 .

(a)

(b)

(c)

(d)

(a)

(b)

(c)

(d)

4.2. Analysis of Simulation Results

(1)Analysis of advance abutment pressure distribution under different rates of advance: Figure 5 shows the distribution diagram of abutment pressure on the advancing (by 100 m) mining face under different rates of advance. Combined with the data in Figure 4, the characteristics of roof pressure at different rates of advance are plotted, using data in Table 5. The analysis shows that, with the continuous increase of the rate of advance, the bearing pressure first decreases and then decelerates. When the rate of advance of the working face reaches 15 m/d, the weakening trend of the advanced abutment pressure will diminish. Compared with the rate of advance of 20 m/d, the stress concentration in each case is similar. Compared with the slow rate of advance, the stress concentration is smaller when the rate of advance reaches 15 m/d; at the same time, it is concluded that, in a certain range of rate of advance, increasing the rate of advance reduces the working face pressure, but when the rate of advance reaches a certain value, the change in the rate of advance has little influence on the weighting of the working face.(2)Analysis of advance abutment pressure distribution in different times of stopping mining under different rates of advance: Taking survey line 1 as an example, the difference in abutment pressure at different stopping times under different rates of advance is analysed. Figure 6 shows the distribution of advanced abutment pressure at different stopping times at different rates of advance of line 1, and Table 6 shows the concentrated characteristics of stopping production pressure at different rates of advance of line 1. Accordingly, the peak position of advanced abutment pressure appears at about 12 m from the coal wall, and the abutment pressure first increases and then decreases with the distance of the leading coal wall; at the same time, the degree of concentration of abutment pressure also increases with the extension of the stoppage time. According to the data in Table 6, the range of pressure change is not large when the mining is stopped for 2 days or 3 days, so it is speculated that the roof pressure concentration had stabilised.(3)Analysis of maximum eigenvalue variation of abutment pressure at different rates of advance: Figure 7 shows the variation in the maximum eigenvalue of different rates of advance. It can be seen from Figure 6 that the maximum eigenvalue gradually decreases with the increase of the rate of advance, and the increment of the stress peak value changes to a significant extent in the process of the gradual acceleration of the rate of advance. When the rate of advance reaches 15 m/d, the peak stress increase presents a downward trend. Compared with the four curves in Figure 7, it can be seen that, with the continuous increase of the rate of advance of the working face, the increase of the maximum eigenvalue is significant: the rate of change of abutment pressure increment at different stoppage times is also accelerated. When the working face stops mining for 1 day, the increment of abutment pressure changes most, then, with the increase of stoppage time, the increment of abutment pressure gradually decreases and finally diminishes; therefore, it can be concluded that the faster the rate of advance of working face, the greater the roof pressure caused by stopping mining. When the rate of advance reaches a certain value, the roof pressure will increase and the increment of abutment pressure then will decrease.

5. Engineering Application

Combining with the characteristics of support load distribution in the 620 working face with different rates of advance in the past, it is necessary to take some pressure relief measures according to the pressure accumulation phenomenon when stopping mining at different rates of advance in the 620 working face, so as to release the roof pressure and maintain the safety and stability of the working face. Taking the stoppage of 620 working face on 8 August 2019 as an example, the current daily rate of advance is 12.8 m/d, and the method of constructing a caving hole is adopted to release roof pressure locally to assist goaf caving. The MQT-120T pneumatic drill pipe machine is used for construction. The first row is 3 m from the coal wall of the goaf, with 13–15 rows in each row. The spacing between rows is 300 × 4000 mm, the hole depth is 8 m, and the hole diameter is 28 mm. At the same time, a discharge top hole is constructed at a spacing of 300 mm from the side slope of the solid coal (Figure 8).

As shown in Figure 9, the two instances of the sudden drop of support load to zero are due to the decline of working face support and the separation of support top beam and roof, resulting in no load above the support. Periodic weighting occurs at the working face at about 06 : 00, and the peak load on the support is 36.2 MPa. The roof then collapses and the support load decreases. The time from peak weighting to stoppage is about 6 h. During this period of time, at about 11 : 00, the top beam of the support dropped and separated from the roof, resulting in a sudden drop of the support load to zero, and at 15 : 07, the working face suffers from tile failure. At this time, the support load increases slowly, the peak load is 23.8 MPa, the load increase is 6.3 MPa, and the duration of pressurisation is 1.8 h. At the same time, forced caving measures are taken to release the roof pressure. At this time, the support load drops sharply, the final load decreases to 11.7 MPa, and the cumulative load decreases by 12.1 MPa. Compared with the support load before and after pressure relief, the load rise rate caused by stopping mining is 3.5 MPa/h, that caused by forced caving is 23.4 MPa/h, the load drop rate is 6.68 times the previous rate of rise, and the forced caving pressure relief effect is obvious.

6. Conclusion

(1)The load transfer velocity factor of a working face is affected by lithology factor K_r and the periodic weighting step. After the working face is calculated, is determined accordingly; that is, the load transfer speed factor applicable to the working face is positively correlated with the rate of advance of the working face.(2)The reasonable hanging roof length based on the allowable support resistance of the support is mainly affected by the number of caving holes and the pressure relief effect.(3)With the continuous increase of the rate of advance of working face, the advance abutment pressure first decreases and then tends to be stable. At different stoppage times, the incremental trend of roof abutment pressure varies: with the continuous extension of stoppage time, the increment of roof abutment pressure decreases, but when the rate of advance increases to within a certain range, the trend of abutment pressure increment decreases.(4)The roof has accumulated part of energy after the working face has been excavated continuously after normal cycle weighting. If a stoppage then occurs, the roof pressure accumulation is significant, so it is necessary to take forced roof caving measures to control the roof pressure.

Data Availability

The data used to support the findings of this study are available from the corresponding author upon request.

Conflicts of Interest

The authors declare that they have no conflicts of interest.

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Copyright © 2021 Xiaoxu Gao 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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