# Floor Failure Evolution Mechanism for a Fully Mechanized Longwall Mining Face above a Confined Aquifer

Authors: Jinghui Zhai; Danlong Liu; Gang Li; Fangtian Wang
Journal: Advances in Civil Engineering
 (2019)

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## Abstract

In longwall mining, the risk of water inrushes from the floors of deeply buried coal seams is closely related to the degree and depth of the destruction for the mining floor. To analyze the main factors affecting floor failure and the evolution of such failures, this study considered the LW2703 working face of the Chengjiao Coal Mine in China, which is characterized by a large buried depth, complex fault structure, and high pressure from a confined aquifer. The characteristics affecting floor crack development depth were analyzed by considering friction angle, cohesion force, floor pressure, stress increase coefficient, and peak position. A FLAC3D simulation was performed to compare the degrees of floor damage that occurred for caving and backfilling methods during the mining process. High-density electrical detection was performed on-site and used to (1) determine the maximum depth range of the floor damage, (2) reveal the laws governing the evolution of damage in a mining floor, and (3) provide a reasonable basis for evaluating and preventing floor water inrush accidents.


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## Body

## 1. Introduction

As Chinese coal mines gradually become deeper, their hydrogeological conditions become increasingly complex. The combination of ground and water pressure has made water damage more prominent and increased the frequency of floor water inrush accidents. Heavy casualties and economic losses, combined with high-frequency growth, have become a critical risk to the safe and efficient operation of coal mines [1, 2]. Figure 1 shows that while water hazard accidents and deaths in coal mines have generally decreased in recent years in China, heavy casualties and economic losses still occur. Therefore, further investigation is needed into the detection and prevention of water hazards in mines.Figure 1
Coal mine water disasters in China between 2000 and 2017.Floor water inrushing occurs when certain basic conditions are satisfied for a water source and channel. For the north China coalfield, the water source is generally the water-filled aquifer of the Carboniferous and Ordovician systems. This indicates that the safety of a mining face depends mainly on channel formation, which is controlled by many factors. Hidden structure activation leads to water inrushing [3], which indicates that water damage is closely related to the degree and depth of structural damage. Therefore, studying the evolution of mining floor failures is of great theoretical and practical significance.Many researchers in China and abroad have investigated mining floor failure laws and water inrush controls. Yao et al. [4] presented a suite of fully coupled governing equations to determine fracture erosion and changes in rock permeability. To determine the hydraulic conductivity of a rock formation between a deeply buried coal seam and an aquifer, Huang et al. [5] conducted four water injection tests at the Baodian Coal Mine. Zhai et al. [6] simulated and analyzed the failure characteristics at different depths under fluid-solid coupling conditions to develop a water inrush law. Li et al. [7] used a double scalar D-P elastoplastic damage constitutive model to study the evolution of a complete base plate from an aquifer to an aqueduct. Wang [8] analyzed key methods for preventing floor water damage in the Fengfeng mining area and proposed comprehensive prevention and control techniques, such as geophysical exploration and ground grouting reinforcement. Wu et al. [9] systematically classified mine water hazards and provided a basis for their classification and prevention. Chen and Cao [10] analyzed the hydrogeological conditions and karst development rules for karst water in the Pingdingshan Coal Mine, as well as the water-filling conditions of the mine and its karst fissure distribution. Wang et al. [11, 12] performed a Hopkinson pressure bar experiment to study the dynamic failure characteristics of coal and indicated that its dynamic compressive strength was positively correlated with its elastic modulus. Guo et al. [13] divided flooding patterns into three types: complete floor crack extension, primary channel conduction, and concealed structure sliding shear. They studied different modes for the spatiotemporal evolution of water inrush channels.The LW2703 working face in the Chengjiao Coal Mine is very deep, with a complex fault structure and high water pressure. Because of these characteristics, the face has a high water inrush risk, which threatens safe mining and production. We investigated the LW2703 working face and analyzed how floor crack development depth was affected by friction angle, cohesion force, floor pressure, stress increase coefficients, and peak position. We used a FLAC3D simulation to compare and analyze how caving and backfilling methods affected the degree of floor damage during mining. High-density electrical detection was performed in the field and used to (1) determine the maximum depth of floor damage, (2) reveal the laws governing the evolution of damage in a mining floor, and (3) provide a reasonable basis for evaluating and preventing floor water inrush accidents.


## 2. Engineering Background

The LW2703 working face of the Chengjiao Coal Mine is located in the seventh mining area of the east wing. The working face has a length of 1043 m along the upper roadway, 926 m along the lower roadway, and 993 m along the middle roadway. The inclined length of the working face is 342.2 m. Figure2 shows the layout of the working face. The actual exposure during roadway excavation showed that the No. 2 coal seams are stable, with an average coal thickness of 2.8 m, buried depth of over 800 m, and inclination angle of 5°. Figure 3 shows the roof and floor. The floor consists of sandstone and mudstone. The limestone layers below the floor (L8, L9, L10, and L11) are rich in aquifers and have low water conductivity, and the cracks in the floor are undeveloped. Instead, the direct water source for the mine is the sandstone in the roof and floor. The indirect water supply source is the limestone in L11 and L10. The main water source that threatens the working face is the upper limestone of the Taiyuan Formation. The L11 and L10 limestone layers are located 74.0 and 63.5 m below the No. 2 coal seam, respectively. The water pressure is 4.2–4.5 MPa. The limestone and No. 2 coal seam have relatively stable mudstone and sandy mudstone phase barriers, and water inrushing is unlikely to occur under normal conditions. However, Mining activities can cause floor damage and crack the floor strata. At the same time, the natural fissures in the lower part of the aquifer develop upwards, which further reduces the thickness of the water-resisting layer and can cause water inrush accidents at the working face.Figure 2
LW2703 working face layout.Figure 3
Coal seam histogram.


## 3. Main Factors Affecting the Degree of Floor Damage

### 3.1. Factors Affecting Floor Crack Development Depth

The rock strength index has a certain effect on the development of floor cracks. In this work, we analyzed the effects of the floor crack development depth on the internal friction angle, cohesion force, floor pressure, stress increasing coefficient, and peak position were analyzed. The shear stress near the coal wall was high, a condition that can easily cause fissures. The crack depth reached a maximum value within 150 m from the cutting eye and 30 m from the coal wall. Therefore, this zone was selected for the tests and analysis.


#### 3.1.1. Relationship between Internal Friction Angle, Cohesion, and Floor Crack Development Depth

Figure4 shows fitting curves for the internal friction angle and cohesion, based on tests.Figure 4
Fitting curves of fracture development depth versus (a) internal friction angle and (b) cohesion.
(a)
(b)Figure4(a) shows a positive correlation, which indicates that a larger internal friction angle corresponds to a larger floor crack development depth and vice versa. The two parameters have a power exponential relationship, which can be expressed as follows:(1)h=3.0926e0.2608a,R2=0.9819,where α is the internal friction angle in the rock body and R2 is the reliability.Figure4(b) shows a linear relationship between crack development depth and cohesion. Crack development depth increases with cohesion, and this linear relationship can be represented as follows:(2)h=0.2681C2−0.1742C,R2=0.09963,where C is the cohesion in the rock body (MPa).


#### 3.1.2. Relationship between the Floor Water Pressure and Depth of Floor Crack Development

Changes in the pressure of the confined water influence floor cracks. This can be expressed by the curves shown in Figure5.Figure 5
Curve for water pressure and fracture development.The development depth of the floor cracks increases slowly between confined water pressures of 0 to 2 MPa and then increases rapidly for higher pressures. The floor crack development depths were 10.8 and 17 m at 2 and 3 MPa, respectively, which represents an increase of 57.4%. The depth at 4 MPa was 24 m. The relationship is expressed as follows:(3)h=0.9571p2−1.2829p+6.54,R2=0.9976,where P is the water pressure and R2 is the reliability.


#### 3.1.3. Relationship between Abutment Pressure Concentration Coefficient and Depth of Cracks in the Floor

The floor crack development depth changes with the bearing pressure concentration coefficient, as shown by the curve in Figure6.Figure 6
Curve of the accommodating pressure concentration and crack development.The fissure development depth first decreases and then increases with the concentration coefficient. The concentration coefficient reaches a minimum between 2.5 and 3.5 and is approximately 3, which indicates that the floor crack development depth is shallow. In addition, intersection points 2 and 4 for the actual and fitted curves indicate that the two parameters have an equal effect when the lumping factors are 2 and 4. The linear relationship is expressed as follows:(4)h=1.525k2−7.563k+17.85,R2=0.8688,where K is the concentration pressure coefficient.


#### 3.1.4. Relationship between the Peak Position of the Support Pressure and Crack Depth

The continuous advancement of the coal mining face means that the distance between the working surface and peak position of the support pressure is constantly changing. This results in a corresponding change in the crack depth. Figure7 shows the curve for this relationship.Figure 7
Fitting curve of the crack development depth and support pressure at the floor.As the mining face continues to advance, the distance between the location of the peak bearing pressure and the mining face increases, and the floor crack development depth slowly increases before reaching a plateau. At this point, increasing the distance no longer increases the damage and depth. The relationship is expressed as follows:(5)h=1.1199lnL+6.7277,R2=0.9648,where L is the distance between the mining face and the location of the peak abutment pressure.


#### 3.1.5. Floor Damage Depth Induced by Longwall Mining

As shown in Figure8, the maximum floor damage depth ha induced by longwall mining can be calculated with fracture mechanics theory.Figure 8
Depth of the mining failure of the floor.The working face is assumed to be a crack in the internal part of an infinite rock. For the working face, the mining thickness is much smaller than the mining width. Consequently, the maximum damage depthha in the floor can be calculated as follows [14]:(6)ha=1.57γ2H2a1−sinφ216c2cos2φ,where γ is the average weight of the overlying coal seam layer (25 kN/m3), H is the buried depth of the coal seam (846 m), c and φ are the average cohesion (15.8 MPa) and internal friction angle (30°), respectively, for the floor rock, and a is the width of the working face (342.2 m). As a result, the floor damage depth was calculated to be 20.1 m.


### 3.2. Simulation Analysis of the Floor Damage Degree

FLAC3D was used to develop the numerical models for caving and backfilling mining.


#### 3.2.1. Stress Distribution of the Mining Floor

Figure9 shows the vertical stress contours 2 m below the floor when the working face was advanced by 80, 120, and 160 m. When the roof was controlled with a caving method, the stress in the front and on both sides of the working face increased, and stress in the goaf area decreased. The area of increased stress advanced with the work surface. The maximum stress in front of the working face was 8.5 MPa, which was 1.5 times that of the original rock, and the bearing pressure affected the rock within a range of 35–40 m. Both sides of the working face appear on the change of stress as shown in Figure 10. The vertical stress on the rock formations decreased. The stress in the goaf is gradually recovering as the working face advanced.Figure 9
Vertical stress of the floor at different advance distances: 80 m mining advance by (a) caving and (b) backfilling; 120 m mining advance by (c) caving and (d) backfilling; and 160 m mining advance by (e) caving and (f) backfilling.
(a)
(b)
(c)
(d)
(e)
(f)Figure 10
Stress perspective for a 120 m mining advance: (a) caving and (b) backfilling.
(a)
(b)When the backfilling method was used, stress concentrations also occurred. The maximum stress was 7.2 MPa, which was 1.2 times the stress of the normal rock. The supporting pressure had a range of 25–30 m, which subsequently decreased as the working face advanced. Rock formations were noted in the goaf area. The reduction in vertical stress was lower than the corresponding reduction for the caving method. This indicates that the backfill transferred part of the stress in the goaf, which gradually changed the surrounding rock stress.The overall stress distribution characteristics were similar for the caving and backfilling methods; however, the caving method produced a higher degree of stress concentration, while the backfilling method produced a gentler curve for the change in stress.


#### 3.2.2. Distribution of the Plastic Zone in the Mining Floor

Figure11 shows the distribution of the floor’s plastic zone under the midsection of the working face along the strike. As the plastic zone continued to advance, the floor first entered shear plastic yielding and then experienced tensile yield failure. Based on the mine pressure, the depth of the effects can be divided into three zones: direct damage, impact, and minor change.Figure 11
Development range of the plastic zone for the floor at different advance distances: 80 m mining advance by (a) caving and (b) backfilling; 120 m mining advance by (c) caving and (d) backfilling; and 160 m mining advance by (e) caving and (f) backfilling.
(a)
(b)
(c)
(d)
(e)
(f)For the caving method, when the working face advanced 80 m, shear plastic yielding developed to 24 m below the floor, and tensile yielding developed to 6 m. When the working face advanced 80 m, shear yielding still developed to 24 m, but the range increased; the tensile yielding developed to 7 m below the floor, and its range also expanded. When the working face advanced 120 m, the shear yielding did not change, and the range was wider; the tensile yielding developed to 12 m below the floor and no longer developed downward. The range also continued to expand.For the backfilling method, damage to the floor was greatly reduced, and tensile yielding only occurred around 2 m near the roof and floor. When the working face advanced 80 m, shear yielding developed to 8 m below the floor. When the working face advanced 160 m, shear yielding developed to 11 m. A comparison of the caving and backfilling methods shows that the backfilling method can effectively control failure because the shear yielding depth for backfilling was approximately one-third that for caving.In summary, different mining techniques caused the stress field and plastic zone to have different development ranges. The caving method causes a greater crack development depth than the filling method and more intense damage to the floor.


## 3.1. Factors Affecting Floor Crack Development Depth

The rock strength index has a certain effect on the development of floor cracks. In this work, we analyzed the effects of the floor crack development depth on the internal friction angle, cohesion force, floor pressure, stress increasing coefficient, and peak position were analyzed. The shear stress near the coal wall was high, a condition that can easily cause fissures. The crack depth reached a maximum value within 150 m from the cutting eye and 30 m from the coal wall. Therefore, this zone was selected for the tests and analysis.


### 3.1.1. Relationship between Internal Friction Angle, Cohesion, and Floor Crack Development Depth

Figure4 shows fitting curves for the internal friction angle and cohesion, based on tests.Figure 4
Fitting curves of fracture development depth versus (a) internal friction angle and (b) cohesion.
(a)
(b)Figure4(a) shows a positive correlation, which indicates that a larger internal friction angle corresponds to a larger floor crack development depth and vice versa. The two parameters have a power exponential relationship, which can be expressed as follows:(1)h=3.0926e0.2608a,R2=0.9819,where α is the internal friction angle in the rock body and R2 is the reliability.Figure4(b) shows a linear relationship between crack development depth and cohesion. Crack development depth increases with cohesion, and this linear relationship can be represented as follows:(2)h=0.2681C2−0.1742C,R2=0.09963,where C is the cohesion in the rock body (MPa).


### 3.1.2. Relationship between the Floor Water Pressure and Depth of Floor Crack Development

Changes in the pressure of the confined water influence floor cracks. This can be expressed by the curves shown in Figure5.Figure 5
Curve for water pressure and fracture development.The development depth of the floor cracks increases slowly between confined water pressures of 0 to 2 MPa and then increases rapidly for higher pressures. The floor crack development depths were 10.8 and 17 m at 2 and 3 MPa, respectively, which represents an increase of 57.4%. The depth at 4 MPa was 24 m. The relationship is expressed as follows:(3)h=0.9571p2−1.2829p+6.54,R2=0.9976,where P is the water pressure and R2 is the reliability.


### 3.1.3. Relationship between Abutment Pressure Concentration Coefficient and Depth of Cracks in the Floor

The floor crack development depth changes with the bearing pressure concentration coefficient, as shown by the curve in Figure6.Figure 6
Curve of the accommodating pressure concentration and crack development.The fissure development depth first decreases and then increases with the concentration coefficient. The concentration coefficient reaches a minimum between 2.5 and 3.5 and is approximately 3, which indicates that the floor crack development depth is shallow. In addition, intersection points 2 and 4 for the actual and fitted curves indicate that the two parameters have an equal effect when the lumping factors are 2 and 4. The linear relationship is expressed as follows:(4)h=1.525k2−7.563k+17.85,R2=0.8688,where K is the concentration pressure coefficient.


### 3.1.4. Relationship between the Peak Position of the Support Pressure and Crack Depth

The continuous advancement of the coal mining face means that the distance between the working surface and peak position of the support pressure is constantly changing. This results in a corresponding change in the crack depth. Figure7 shows the curve for this relationship.Figure 7
Fitting curve of the crack development depth and support pressure at the floor.As the mining face continues to advance, the distance between the location of the peak bearing pressure and the mining face increases, and the floor crack development depth slowly increases before reaching a plateau. At this point, increasing the distance no longer increases the damage and depth. The relationship is expressed as follows:(5)h=1.1199lnL+6.7277,R2=0.9648,where L is the distance between the mining face and the location of the peak abutment pressure.


### 3.1.5. Floor Damage Depth Induced by Longwall Mining

As shown in Figure8, the maximum floor damage depth ha induced by longwall mining can be calculated with fracture mechanics theory.Figure 8
Depth of the mining failure of the floor.The working face is assumed to be a crack in the internal part of an infinite rock. For the working face, the mining thickness is much smaller than the mining width. Consequently, the maximum damage depthha in the floor can be calculated as follows [14]:(6)ha=1.57γ2H2a1−sinφ216c2cos2φ,where γ is the average weight of the overlying coal seam layer (25 kN/m3), H is the buried depth of the coal seam (846 m), c and φ are the average cohesion (15.8 MPa) and internal friction angle (30°), respectively, for the floor rock, and a is the width of the working face (342.2 m). As a result, the floor damage depth was calculated to be 20.1 m.


## 3.1.1. Relationship between Internal Friction Angle, Cohesion, and Floor Crack Development Depth

Figure4 shows fitting curves for the internal friction angle and cohesion, based on tests.Figure 4
Fitting curves of fracture development depth versus (a) internal friction angle and (b) cohesion.
(a)
(b)Figure4(a) shows a positive correlation, which indicates that a larger internal friction angle corresponds to a larger floor crack development depth and vice versa. The two parameters have a power exponential relationship, which can be expressed as follows:(1)h=3.0926e0.2608a,R2=0.9819,where α is the internal friction angle in the rock body and R2 is the reliability.Figure4(b) shows a linear relationship between crack development depth and cohesion. Crack development depth increases with cohesion, and this linear relationship can be represented as follows:(2)h=0.2681C2−0.1742C,R2=0.09963,where C is the cohesion in the rock body (MPa).


## 3.1.2. Relationship between the Floor Water Pressure and Depth of Floor Crack Development

Changes in the pressure of the confined water influence floor cracks. This can be expressed by the curves shown in Figure5.Figure 5
Curve for water pressure and fracture development.The development depth of the floor cracks increases slowly between confined water pressures of 0 to 2 MPa and then increases rapidly for higher pressures. The floor crack development depths were 10.8 and 17 m at 2 and 3 MPa, respectively, which represents an increase of 57.4%. The depth at 4 MPa was 24 m. The relationship is expressed as follows:(3)h=0.9571p2−1.2829p+6.54,R2=0.9976,where P is the water pressure and R2 is the reliability.


## 3.1.3. Relationship between Abutment Pressure Concentration Coefficient and Depth of Cracks in the Floor

The floor crack development depth changes with the bearing pressure concentration coefficient, as shown by the curve in Figure6.Figure 6
Curve of the accommodating pressure concentration and crack development.The fissure development depth first decreases and then increases with the concentration coefficient. The concentration coefficient reaches a minimum between 2.5 and 3.5 and is approximately 3, which indicates that the floor crack development depth is shallow. In addition, intersection points 2 and 4 for the actual and fitted curves indicate that the two parameters have an equal effect when the lumping factors are 2 and 4. The linear relationship is expressed as follows:(4)h=1.525k2−7.563k+17.85,R2=0.8688,where K is the concentration pressure coefficient.


## 3.1.4. Relationship between the Peak Position of the Support Pressure and Crack Depth

The continuous advancement of the coal mining face means that the distance between the working surface and peak position of the support pressure is constantly changing. This results in a corresponding change in the crack depth. Figure7 shows the curve for this relationship.Figure 7
Fitting curve of the crack development depth and support pressure at the floor.As the mining face continues to advance, the distance between the location of the peak bearing pressure and the mining face increases, and the floor crack development depth slowly increases before reaching a plateau. At this point, increasing the distance no longer increases the damage and depth. The relationship is expressed as follows:(5)h=1.1199lnL+6.7277,R2=0.9648,where L is the distance between the mining face and the location of the peak abutment pressure.


## 3.1.5. Floor Damage Depth Induced by Longwall Mining

As shown in Figure8, the maximum floor damage depth ha induced by longwall mining can be calculated with fracture mechanics theory.Figure 8
Depth of the mining failure of the floor.The working face is assumed to be a crack in the internal part of an infinite rock. For the working face, the mining thickness is much smaller than the mining width. Consequently, the maximum damage depthha in the floor can be calculated as follows [14]:(6)ha=1.57γ2H2a1−sinφ216c2cos2φ,where γ is the average weight of the overlying coal seam layer (25 kN/m3), H is the buried depth of the coal seam (846 m), c and φ are the average cohesion (15.8 MPa) and internal friction angle (30°), respectively, for the floor rock, and a is the width of the working face (342.2 m). As a result, the floor damage depth was calculated to be 20.1 m.


## 3.2. Simulation Analysis of the Floor Damage Degree

FLAC3D was used to develop the numerical models for caving and backfilling mining.


### 3.2.1. Stress Distribution of the Mining Floor

Figure9 shows the vertical stress contours 2 m below the floor when the working face was advanced by 80, 120, and 160 m. When the roof was controlled with a caving method, the stress in the front and on both sides of the working face increased, and stress in the goaf area decreased. The area of increased stress advanced with the work surface. The maximum stress in front of the working face was 8.5 MPa, which was 1.5 times that of the original rock, and the bearing pressure affected the rock within a range of 35–40 m. Both sides of the working face appear on the change of stress as shown in Figure 10. The vertical stress on the rock formations decreased. The stress in the goaf is gradually recovering as the working face advanced.Figure 9
Vertical stress of the floor at different advance distances: 80 m mining advance by (a) caving and (b) backfilling; 120 m mining advance by (c) caving and (d) backfilling; and 160 m mining advance by (e) caving and (f) backfilling.
(a)
(b)
(c)
(d)
(e)
(f)Figure 10
Stress perspective for a 120 m mining advance: (a) caving and (b) backfilling.
(a)
(b)When the backfilling method was used, stress concentrations also occurred. The maximum stress was 7.2 MPa, which was 1.2 times the stress of the normal rock. The supporting pressure had a range of 25–30 m, which subsequently decreased as the working face advanced. Rock formations were noted in the goaf area. The reduction in vertical stress was lower than the corresponding reduction for the caving method. This indicates that the backfill transferred part of the stress in the goaf, which gradually changed the surrounding rock stress.The overall stress distribution characteristics were similar for the caving and backfilling methods; however, the caving method produced a higher degree of stress concentration, while the backfilling method produced a gentler curve for the change in stress.


### 3.2.2. Distribution of the Plastic Zone in the Mining Floor

Figure11 shows the distribution of the floor’s plastic zone under the midsection of the working face along the strike. As the plastic zone continued to advance, the floor first entered shear plastic yielding and then experienced tensile yield failure. Based on the mine pressure, the depth of the effects can be divided into three zones: direct damage, impact, and minor change.Figure 11
Development range of the plastic zone for the floor at different advance distances: 80 m mining advance by (a) caving and (b) backfilling; 120 m mining advance by (c) caving and (d) backfilling; and 160 m mining advance by (e) caving and (f) backfilling.
(a)
(b)
(c)
(d)
(e)
(f)For the caving method, when the working face advanced 80 m, shear plastic yielding developed to 24 m below the floor, and tensile yielding developed to 6 m. When the working face advanced 80 m, shear yielding still developed to 24 m, but the range increased; the tensile yielding developed to 7 m below the floor, and its range also expanded. When the working face advanced 120 m, the shear yielding did not change, and the range was wider; the tensile yielding developed to 12 m below the floor and no longer developed downward. The range also continued to expand.For the backfilling method, damage to the floor was greatly reduced, and tensile yielding only occurred around 2 m near the roof and floor. When the working face advanced 80 m, shear yielding developed to 8 m below the floor. When the working face advanced 160 m, shear yielding developed to 11 m. A comparison of the caving and backfilling methods shows that the backfilling method can effectively control failure because the shear yielding depth for backfilling was approximately one-third that for caving.In summary, different mining techniques caused the stress field and plastic zone to have different development ranges. The caving method causes a greater crack development depth than the filling method and more intense damage to the floor.


## 3.2.1. Stress Distribution of the Mining Floor

Figure9 shows the vertical stress contours 2 m below the floor when the working face was advanced by 80, 120, and 160 m. When the roof was controlled with a caving method, the stress in the front and on both sides of the working face increased, and stress in the goaf area decreased. The area of increased stress advanced with the work surface. The maximum stress in front of the working face was 8.5 MPa, which was 1.5 times that of the original rock, and the bearing pressure affected the rock within a range of 35–40 m. Both sides of the working face appear on the change of stress as shown in Figure 10. The vertical stress on the rock formations decreased. The stress in the goaf is gradually recovering as the working face advanced.Figure 9
Vertical stress of the floor at different advance distances: 80 m mining advance by (a) caving and (b) backfilling; 120 m mining advance by (c) caving and (d) backfilling; and 160 m mining advance by (e) caving and (f) backfilling.
(a)
(b)
(c)
(d)
(e)
(f)Figure 10
Stress perspective for a 120 m mining advance: (a) caving and (b) backfilling.
(a)
(b)When the backfilling method was used, stress concentrations also occurred. The maximum stress was 7.2 MPa, which was 1.2 times the stress of the normal rock. The supporting pressure had a range of 25–30 m, which subsequently decreased as the working face advanced. Rock formations were noted in the goaf area. The reduction in vertical stress was lower than the corresponding reduction for the caving method. This indicates that the backfill transferred part of the stress in the goaf, which gradually changed the surrounding rock stress.The overall stress distribution characteristics were similar for the caving and backfilling methods; however, the caving method produced a higher degree of stress concentration, while the backfilling method produced a gentler curve for the change in stress.


## 3.2.2. Distribution of the Plastic Zone in the Mining Floor

Figure11 shows the distribution of the floor’s plastic zone under the midsection of the working face along the strike. As the plastic zone continued to advance, the floor first entered shear plastic yielding and then experienced tensile yield failure. Based on the mine pressure, the depth of the effects can be divided into three zones: direct damage, impact, and minor change.Figure 11
Development range of the plastic zone for the floor at different advance distances: 80 m mining advance by (a) caving and (b) backfilling; 120 m mining advance by (c) caving and (d) backfilling; and 160 m mining advance by (e) caving and (f) backfilling.
(a)
(b)
(c)
(d)
(e)
(f)For the caving method, when the working face advanced 80 m, shear plastic yielding developed to 24 m below the floor, and tensile yielding developed to 6 m. When the working face advanced 80 m, shear yielding still developed to 24 m, but the range increased; the tensile yielding developed to 7 m below the floor, and its range also expanded. When the working face advanced 120 m, the shear yielding did not change, and the range was wider; the tensile yielding developed to 12 m below the floor and no longer developed downward. The range also continued to expand.For the backfilling method, damage to the floor was greatly reduced, and tensile yielding only occurred around 2 m near the roof and floor. When the working face advanced 80 m, shear yielding developed to 8 m below the floor. When the working face advanced 160 m, shear yielding developed to 11 m. A comparison of the caving and backfilling methods shows that the backfilling method can effectively control failure because the shear yielding depth for backfilling was approximately one-third that for caving.In summary, different mining techniques caused the stress field and plastic zone to have different development ranges. The caving method causes a greater crack development depth than the filling method and more intense damage to the floor.


## 4. Exploring the Evolution of Mining-Induced Floor Failures

Compared with traditional drilling methods, high-density resistivity imaging technology is advantageous because it can be used efficiently at construction sites; it has a large detection range and can be used to perform continuous and dynamic observations. It is used to obtain the electrical occurrence state of coal seams before and after mining and to intuitively analyze and judge the extent of damage depth of mining. The continuous detection capabilities of high-density resistivity imaging were used to dynamically monitor the LW2703 working face to study the damage rules and failure depth of the working face floor.We adopted a system using the WDJD-3 high-density electric method. This system has a large storage capacity, performs accurate and fast measurements, is convenient to operate, and is easy to use with domestic high-density electrical processing software, which makes interpretation more convenient and intuitive.


### 4.1. Drilling Arrangement

Borehole #1 was drilled at 713 m along the LW2703 middle roadway, with a drilling orientation of 67°, a drilling inclination of −35°, and a drilling length of 120 m. Borehole #2 was drilled at 640 m along the middle roadway of LW2703, with the same drilling parameters used for borehole #1. The layout is shown in Figure2.


### 4.2. Results Analysis of the Borehole #1 Resistivity

Figure12 shows the resistivity detection results for borehole #1 at different locations. Changes in the apparent resistivity are indicated by colors; red indicates high apparent resistivity and blue indicates low apparent resistivity. The thick blue dotted line shows the difference relative to the previous detection. A total of 29 electrodes were used to perform measurements; the electrodes were spaced 4 m apart. A-MN-B (α) and MN-B devices were used for detections. Comparative tests showed that the α device had a large detection area, good stability, and high sensitivity to subtle electrical changes. The detection results for the α device are explained in the remainder of this section. Table 1 presents the parameters for borehole #1 at different stopping locations.Figure 12
Resistivity results for borehole #1 at different locations on the working face: (a) 814 m, (b) 807 m, (c) 802 m, (d) 796 m, (e) 793 m, (f) 783 m, (g) 762 m, (h) 682 m, and (i) 557 m.
(a)
(b)
(c)
(d)
(e)
(f)
(g)
(h)
(i)Table 1
List of detection parameters.
No.
Mining position (m)
Detection results
1
814
The electrical characteristics are chaotic, and low resistance characteristics appear near the borehole.
2
807
Two obvious electrical changes: (1) a 20–45 m hole depth and a 0–10 m vertical distance to the borehole; (2) a 40–60 m hole depth, a 20–30 m vertical distance to borehole, and a 0–5 m vertical distance to coal seam floor.
3
802
One obvious electrical change: a 45–65 m hole depth, an approximately 15–30 m vertical distance to the borehole, and an approximately 0–14.3 m vertical distance to the coal seam floor.
4
796
Two obvious electrical changes: (1) 5–30 m hole depth, 0–10 m vertical distance to drill hole, and 0–14 m vertical distance to floor; (2) 50–70 m hole depth, 15–35 m vertical distance to drill hole, and about 0–17 m vertical distance to coal floor.
5
793
One obvious electrical change: a 45–65 m hole depth, a 20–40 m vertical distance to the drill hole, and a 0–5.2 m vertical distance to the floor.
6
783
Two obvious electrical changes: (1) at a hole depth of 45–55 m, the location of the work water and the greater grounding of the metal body led to a significant reduction; (2) at a 60–75 m hole depth, resistivity was significantly higher than that previously measured. The resistance value suddenly changes from 100 Ω·m to 1000 Ω·m. Based on the mining conditions, it is judged that the floor is broken due to the mining effect. There was no water in the initial stage or slight water cutting. The characteristics of the rock formation clearly were highly electrical.
7
762
One obvious electrical change: a 45–80 m hole depth, a 12–40 m vertical distance to the drill hole, and a 0–24.6 m offset distance from the coal floor. The site was judged to be the floor rock layer destroyed by mining.
8
682
One obvious electrical change: a 25–80 m hole depth. The site was judged to be the floor destroyed by mining.
9
557
Within the same range as the previous change, the rock formation in the initial stage was severely broken by mining. It gradually became compact and stable. The resistance was generally low. There was a certain degree of water absorption, and the distribution was more uniform.Multiple explorations of the two boreholes, combined with the exploration map and the geological conditions of the working face, revealed the following:(1)
Multiple comparisons showed obvious electrical changes, and the rules governing these changes were in good agreement with the theoretical characteristics. In particular, the results for borehole #1 were relatively stable, and the data were accurate and reliable.(2)
In the early stages of drilling, when the water or slurry in the hole was not solidified, a low-resistance screening effect was clearly detected. As the slurry gradually solidified, electrode grounding gradually improved. The detection data gradually approached the electrical characteristics of the formation.(3)
In the early stages of the rock formation destruction at the working face, the rock layer was broken, the size of cracks increased, resistivity increased sharply, there was high disorder, and the destruction depth was shallow. As the working face continued to advance, the failure depth of the floor rock layer gradually increased. The gradual filling of sandstone pores and fissures with water caused resistivity to decrease sharply. The apparent resistivity fell below 100 Ω·m.(4)
The floor of the detection location was severely broken after mining; it gradually became compact and stable, with low resistance. There was some degree of water absorption, but the distribution rules were relatively uniform, and the shallow part gradually moved from very high to medium resistance with increasing depth to form a wide range of layers of low resistance.(5)
The destruction depth of this exploration mainly focused on borehole #1. The deepest change position occurred at a hole depth of 65 m, with a vertical distance of 6 m from the drill hole and 27.3 m from the coal seam.


## 4.1. Drilling Arrangement

Borehole #1 was drilled at 713 m along the LW2703 middle roadway, with a drilling orientation of 67°, a drilling inclination of −35°, and a drilling length of 120 m. Borehole #2 was drilled at 640 m along the middle roadway of LW2703, with the same drilling parameters used for borehole #1. The layout is shown in Figure2.


## 4.2. Results Analysis of the Borehole #1 Resistivity

Figure12 shows the resistivity detection results for borehole #1 at different locations. Changes in the apparent resistivity are indicated by colors; red indicates high apparent resistivity and blue indicates low apparent resistivity. The thick blue dotted line shows the difference relative to the previous detection. A total of 29 electrodes were used to perform measurements; the electrodes were spaced 4 m apart. A-MN-B (α) and MN-B devices were used for detections. Comparative tests showed that the α device had a large detection area, good stability, and high sensitivity to subtle electrical changes. The detection results for the α device are explained in the remainder of this section. Table 1 presents the parameters for borehole #1 at different stopping locations.Figure 12
Resistivity results for borehole #1 at different locations on the working face: (a) 814 m, (b) 807 m, (c) 802 m, (d) 796 m, (e) 793 m, (f) 783 m, (g) 762 m, (h) 682 m, and (i) 557 m.
(a)
(b)
(c)
(d)
(e)
(f)
(g)
(h)
(i)Table 1
List of detection parameters.
No.
Mining position (m)
Detection results
1
814
The electrical characteristics are chaotic, and low resistance characteristics appear near the borehole.
2
807
Two obvious electrical changes: (1) a 20–45 m hole depth and a 0–10 m vertical distance to the borehole; (2) a 40–60 m hole depth, a 20–30 m vertical distance to borehole, and a 0–5 m vertical distance to coal seam floor.
3
802
One obvious electrical change: a 45–65 m hole depth, an approximately 15–30 m vertical distance to the borehole, and an approximately 0–14.3 m vertical distance to the coal seam floor.
4
796
Two obvious electrical changes: (1) 5–30 m hole depth, 0–10 m vertical distance to drill hole, and 0–14 m vertical distance to floor; (2) 50–70 m hole depth, 15–35 m vertical distance to drill hole, and about 0–17 m vertical distance to coal floor.
5
793
One obvious electrical change: a 45–65 m hole depth, a 20–40 m vertical distance to the drill hole, and a 0–5.2 m vertical distance to the floor.
6
783
Two obvious electrical changes: (1) at a hole depth of 45–55 m, the location of the work water and the greater grounding of the metal body led to a significant reduction; (2) at a 60–75 m hole depth, resistivity was significantly higher than that previously measured. The resistance value suddenly changes from 100 Ω·m to 1000 Ω·m. Based on the mining conditions, it is judged that the floor is broken due to the mining effect. There was no water in the initial stage or slight water cutting. The characteristics of the rock formation clearly were highly electrical.
7
762
One obvious electrical change: a 45–80 m hole depth, a 12–40 m vertical distance to the drill hole, and a 0–24.6 m offset distance from the coal floor. The site was judged to be the floor rock layer destroyed by mining.
8
682
One obvious electrical change: a 25–80 m hole depth. The site was judged to be the floor destroyed by mining.
9
557
Within the same range as the previous change, the rock formation in the initial stage was severely broken by mining. It gradually became compact and stable. The resistance was generally low. There was a certain degree of water absorption, and the distribution was more uniform.Multiple explorations of the two boreholes, combined with the exploration map and the geological conditions of the working face, revealed the following:(1)
Multiple comparisons showed obvious electrical changes, and the rules governing these changes were in good agreement with the theoretical characteristics. In particular, the results for borehole #1 were relatively stable, and the data were accurate and reliable.(2)
In the early stages of drilling, when the water or slurry in the hole was not solidified, a low-resistance screening effect was clearly detected. As the slurry gradually solidified, electrode grounding gradually improved. The detection data gradually approached the electrical characteristics of the formation.(3)
In the early stages of the rock formation destruction at the working face, the rock layer was broken, the size of cracks increased, resistivity increased sharply, there was high disorder, and the destruction depth was shallow. As the working face continued to advance, the failure depth of the floor rock layer gradually increased. The gradual filling of sandstone pores and fissures with water caused resistivity to decrease sharply. The apparent resistivity fell below 100 Ω·m.(4)
The floor of the detection location was severely broken after mining; it gradually became compact and stable, with low resistance. There was some degree of water absorption, but the distribution rules were relatively uniform, and the shallow part gradually moved from very high to medium resistance with increasing depth to form a wide range of layers of low resistance.(5)
The destruction depth of this exploration mainly focused on borehole #1. The deepest change position occurred at a hole depth of 65 m, with a vertical distance of 6 m from the drill hole and 27.3 m from the coal seam.


## 5. Conclusions

(1)
The LW2703 working face has a relatively complex geology. The limestone in the upper part of the Taiyuan Formation is the main source of water threatening the safety of the working face. This limestone has a water pressure of 4.2–4.5 MPa.
(2)
The floor damage depth is related to the floor rock formation lithology in terms of the development degree. The floor rock mass strength index was used to analyze how the friction angle, cohesion force, floor pressure, stress increase coefficient, and peak position affected floor crack development depth.
(3)
FLAC3D was used to simulate the stress field and plastic zone development associated with caving and backfilling methods. The caving method produces deeper cracks and more intense floor damage than the filling method.
(4)
High-density electrical surveys indicated that the floor damages and the fractured rock formations occur induced by mining effect. In the initial period, there was no or little water present. The electrical characteristics of the rock formations were significantly stronger, and the maximum depth of the floor damage proved the mining floor failure evolution rule.



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*Source: 8036928-2019-01-13.xml*