# Research on Stability Control Technology of Mining Roadway Based on Energy Transformation

Authors: Guohua Zhang; Yanwei Duan; Tao Qin
Journal: Geofluids
 (2023)

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## Abstract

The process of roadway surrounding rock deformation and instability is always accompanied by energy changes, and the energy transformation of surrounding rock directly drives its deformation and failure. In order to realize the stability control of mining roadway, the distribution of elastic strain energy and plastic strain energy of surrounding rock was analyzed by using FLAC3D numerical simulation software on the basis of the energy transformation law of sandstone under different confining pressures revealed by indoor triaxial compression test. Based on this, combined with the energy transformation, the principle of roadway surrounding rock stability control technology is proposed. One is to reduce the elastic strain energy of surrounding rock, that is, by increasing the extension of support body, part of the roadway energy is transferred to the support body, so as to improve the energy release of surrounding rock, or by optimizing the layout of roadway to reduce the energy accumulation of surrounding rock. The second is to increase the plastic strain energy of surrounding rock, that is, to increase the energy dissipation of surrounding rock by setting weak structure. The above research results are applied to the surrounding rock control of the right second transport roadway in the 91st coal of Xinjian coal mine. After the optimization of support measures, the roadway surface displacement and its increase rate decreased significantly, roof subsidence decreased from 247 mm to 62 mm, floor heave decreased from 120 mm to 26 mm, and the right rib shrinkage decreased from 292 mm to 94 mm, that is, the floor heave control effect was particularly obvious. The deformation of surrounding rock gradually stabilized after 45 days.


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## Body

## 1. Introduction

The rapid development of China’s economy is inseparable from energy. Coal accounts for about 70% of the primary energy consumption structure. Therefore, the safe production of coal mines is of great significance to economic development [1–3] in coal mine safety accidents, the accidents caused by deformation and instability of surrounding rock account for a large proportion [3]. Ensuring the stability of surrounding rock is the premise of coal mine production. In the study of surrounding rock stability, scholars often analyzed the surrounding rock stability from the perspective of mechanics, but the complex geological conditions (large buried depth, faults, soft rock, etc.) often made the mechanical analysis process become complicated [4–8]. In fact, the deformation and failure process of surrounding rock is a process with energy changes from local deformation to overall damage and finally failure. Due to energy accumulation caused by roadway excavation, the energy may be higher than its energy storage limit, which causes deformation of the surrounding rock, and this part of energy will be released concentratedly, resulting in the failure of the surrounding rock. Energy change is the external performance of material physical change, and the great change will cause material damage. Therefore, energy transformation directly drives deformation and failure of surrounding rock, from the perspective of energy interpretation of surrounding rock can reflect its essence [9–11].For the study of surrounding rock stability related to energy, scholars have done a lot of research from laboratory experiments and numerical simulation. On the one hand, laboratory experiments are mainly conducted to explore the characteristics and laws of rock energy evolution through rock mechanics experiments under different stress paths. From the perspective of energy, Xu et al. [12] analyzed the deformation and damage process of sandstone under the action of circulating pore water pressure, and discussed the evolution law of energy absorption and release in this process. Zhao et al. [13] carried out three confining pressure unloading tests under different unloading paths to study the energy evolution characteristics and evolution rate of rock under unloading conditions. By carrying out triaxial cyclic loading and unloading experiments, Yang et al. [14] discussed the energy evolution characteristics of postpeak stress drop stage and postpeak residual stress stage. Liu et al. [15] studied the energy evolution law and linear storage law of red sandstone in the creep process by conducting uniaxial graded creep test and uniaxial cyclic loading and unloading test after creep. Meng et al. [16] carried out triaxial cyclic loading and unloading tests of limestone under six kinds of confining pressures, revealing the confining pressure effect of energy evolution process and distribution law of loaded rock samples.On the other hand, the numerical simulation mainly carries out the energy characteristics of surrounding rock under different working conditions. Hao et al. [17] used FLAC3D to simulate the advance process of working face, analyzed the morphological characteristics and energy distribution characteristics of roadway plastic area within 100 m of the front working face when the advance was 400 m, and revealed the relationship between the morphological distribution and energy of roadway plastic area. By using discrete element simulation software PFC, Yu et al. [18] analyzed the energy and dynamic acoustic emission evolution law of roadway excavation, and discussed the influence of different vertical stresses and lateral pressure coefficients on roadway excavation energy evolution. Shi et al. [19] used FLAC3D to simulate the process of working face, analyzed the energy evolution and release law of surrounding rock in stope, and pointed out that the roof and the front of coal wall were the key areas of energy accumulation or release. Wang et al. [20] used FLAC3D simulation software to analyze the evolution law of elastic energy of the roof, two ribs and driving face under different driving speeds, which was used as the basis for establishing the evolution mechanism of rock burst energy zoning. Huang et al. [21] used FLAC3D to simulate the stress distribution before and after filling gangue in goaf, and then obtained the distribution of elastic strain energy before and after filling. Ma et al. [22] used UDEC to analyze the energy evolution law of roadway surrounding rock under roof of different thickness, and optimized roadway support parameters according to the simulation results, effectively solving the problem of large deformation of roadway.Based on the analysis of the above studies, the study of surrounding rock stability related to energy mainly focuses on the laboratory study of rock energy evolution law and the numerical simulation study of surrounding rock energy characteristics, while the study of roadway surrounding rock stability control based on energy characteristics is relatively insufficient. In this paper, the mining roadway of Xinjian coal mine is taken as the engineering background. Triaxial compression tests were carried out in the laboratory to reveal the energy evolution law of sandstone under different confining pressures. FLAC3D numerical simulation software was used to analyze the distribution law of elastic strain energy and plastic strain energy of roadway surrounding rock. Based on this, combined with the energy transfromation, the principle of roadway surrounding rock stability control technology is proposed. The monitoring results showed that the roadway surface displacement and its increasing speed decreased obviously after the optimization of support measures, which effectively solved the roadway control problem and provided reference for roadway surrounding rock stability control under similar conditions.


## 2. Energy Evolution Characteristics of Rock Deformation and Failure Process

### 2.1. Test Scheme

In order to obtain the characteristics of energy evolution during the deformation and failure of the rock specimens, the conventional triaxial loading test was conducted on the sandstone samples from Xinjian Coal Mine in Qitaihe in China with TOP INDUSTRIE Rock 600-50 automatic servo rock rheometer (see Figure1). The rock specimen was a standard cylinder (50 mm in diameter and 100 mm in height). Axial pressure was applied at a rate of 0.1 mm/min and confining pressure was applied at a rate of 0.05 MPa/s. The predetermined confining pressure was 5 MPa, 10 MPa, 15 MPa, and 20 MPa.Figure 1
TOP INDUSTRIE Rock 600-50 automatic servo rock rheometer.


### 2.2. Energy Calculation Method Based on Test Data

The loading process of the rock specimens is always accompanied by the storage and dissipation of energy. Both of the variables representing the energy storage and dissipation inside rock are elastic energy and dissipated energy. In the experiment, these two variables cannot be recorded in real time. Therefore, they are calculated through stress-strain curves that are easily accessible. Figure2 shows the calculation method. The shaded area denotes the elastic strain energy (Ue) stored by the deformation of the rock specimens, and Eu denotes the unloading modulus of the rock specimens. The dissipated energy (Ud) is the area constrained by the unloading modulus, the stress-strain curve and the horizontal axis.Figure 2
Energy calculation method.According to the first law of thermodinamics, supposing the experiment system is a closed system ignoring the thermal energy dissipation in the experiment, the total energyU generated by the work done by external force is [23–27]
(1)U=Ue+Ud=∫σ1dε1=∑i=0n12ε1i+1−ε1iσ1i+σ1i+1,where Ue is elastic strain energy, Ud is dissipated energy and σ1i, ε1i, σ3i, and ε3i are axial stress, axial strain, circumferential stress, and circumferential strain, respectively.The expression of triaxial compression elastic energy is(2)Ue=12Euσ12+2σ32−2μ2σ1σ3+σ2σ3,where Eu is unloading modulus and μ is about 50% ~60% of the peak strength [28–30].


### 2.3. Characteristic Law of Rock Energy Evolution

It is assumed that the loading system is a closed system with no heat exchange with the external environment. In the loading system, external load input mechanical energy is absorbed by rock specimens, the rock specimens elastic deformation occur in the form of elastic strain energy stored within rock specimens. As the load goes on, the elastic strain energy stored by the rock specimens become less and less, and the energy absorbed at this time is mainly used to produce microcracks inside the rock specimens, namely, the part of energy is dissipated. In order to compare and analyze the law of energy evolution corresponding to different stages of stress-strain curves, axial strain is used as the horizontal axis, axial stress as the left axis, and energy density as the right axis. The energy evolution curves of the conventional triaxial compression test sandstone under different confining pressures are shown in Figure3.Figure 3
Energy evolution curve of sandstone under different confining pressure. (a) Confining pressure 5 MPa. (b) Confining pressure 10 MPa. (c) Confining pressure 15 MPa. (d) Confining pressure 20 MPa.
(a)(b)(c)(d)It can be seen that the energy evolution process can be divided into prepeak stage and postpeak stage according to the stress-strain curve in Figure3 and that different energy evolution curves under different confining pressures have similar characteristics.
(1)
As axial strain increases, total energyU increased exponentially before the axial stress peak and the growth rate decreased slightly after the stress peak. The growth rate is correlated with the properties of stress and strain. The maximum of U varied under different confining pressures. It increased with the increase of confining pressure(2)
Elastic strain energyUe indicates the releasable energy stored in the rock specimens, the evolution law is consistent with the stress-strain curve. In the prepeak stage, the difference between Ue and U was not great. After reaching the peak, Ue dropped rapidly. Also, the difference increased as the axial strain increased. The law of Ue is consistent with the stress-strain curve, which is positively associated with the confining pressure(3)
Dissipated energyUd indicates the energy consumed by the rock specimens during plastic deformation, crack expansion and energy release, which is irreversible. The evolution trend of Ud was slow growth before the peak but rapid growth after the peak. It grew almost exponentially and quickly became a major component of energy. The maximum of Ud varied under different confining pressures and increased with the increase in confining pressure. After the peak, the growth rate of Ud also increased slightly with the increase in confining pressureTo study the transformation of the energy in different stages, the corresponding energy characteristics of each stage are drawn in the bar graphs, and Figure4 shows the energy distribution characteristics of different stages.Figure 4
Energy distribution characteristics at different stages. (a) Prepeak stage. (b) Postpeak stage.
(a)(b)As can be seen in Figure4(a), the energy absorbed in the prepeak stage was mainly converted to elastic strain energy, accounting for about 72% ~80% of the total energy, while dissipated energy remained 18% to 20%. As confining pressure increased, the amount of the stored elastic strain energy became larger. The elastic strain energy was 464 kJ·m-3 when confining pressure was 5 MPa, and when the confining pressure increased to 20 MPa, the elastic strain energy was 1102.6 kJ·m-3, with an apparent confining pressure effect. At this stage, the elastic strain energy generated by rock specimen’s deformation can be continuously stored, while energy is consumed for the closure of cracks inside the rock specimens and the initiation and propagation of microcracks. As can be seen From Figure 4(b), the vast majority of the energy absorbed in the postpeak stage was transferred to dissipated energy, accounting for about 87% of the total energy. As the confining pressure increased, dissipated energy also increased. When confining pressure was 5 MPa, dissipated energy was 787 kJ·m-3. When confining pressure increased to 20 MPa, elastic strain energy was 1733 kJ·m-3. Elastic strain energy only accounted for a very small proportion. In this stage, a large number of macroscopic cracks form in the rock specimens and the total energy is mainly used for the formation and penetration of the macroscopic cracks. However, due to the existence of confining pressure, the rock specimens have certain residual strength, so a small amount of elastic strain energy can still be stored.


## 2.1. Test Scheme

In order to obtain the characteristics of energy evolution during the deformation and failure of the rock specimens, the conventional triaxial loading test was conducted on the sandstone samples from Xinjian Coal Mine in Qitaihe in China with TOP INDUSTRIE Rock 600-50 automatic servo rock rheometer (see Figure1). The rock specimen was a standard cylinder (50 mm in diameter and 100 mm in height). Axial pressure was applied at a rate of 0.1 mm/min and confining pressure was applied at a rate of 0.05 MPa/s. The predetermined confining pressure was 5 MPa, 10 MPa, 15 MPa, and 20 MPa.Figure 1
TOP INDUSTRIE Rock 600-50 automatic servo rock rheometer.


## 2.2. Energy Calculation Method Based on Test Data

The loading process of the rock specimens is always accompanied by the storage and dissipation of energy. Both of the variables representing the energy storage and dissipation inside rock are elastic energy and dissipated energy. In the experiment, these two variables cannot be recorded in real time. Therefore, they are calculated through stress-strain curves that are easily accessible. Figure2 shows the calculation method. The shaded area denotes the elastic strain energy (Ue) stored by the deformation of the rock specimens, and Eu denotes the unloading modulus of the rock specimens. The dissipated energy (Ud) is the area constrained by the unloading modulus, the stress-strain curve and the horizontal axis.Figure 2
Energy calculation method.According to the first law of thermodinamics, supposing the experiment system is a closed system ignoring the thermal energy dissipation in the experiment, the total energyU generated by the work done by external force is [23–27]
(1)U=Ue+Ud=∫σ1dε1=∑i=0n12ε1i+1−ε1iσ1i+σ1i+1,where Ue is elastic strain energy, Ud is dissipated energy and σ1i, ε1i, σ3i, and ε3i are axial stress, axial strain, circumferential stress, and circumferential strain, respectively.The expression of triaxial compression elastic energy is(2)Ue=12Euσ12+2σ32−2μ2σ1σ3+σ2σ3,where Eu is unloading modulus and μ is about 50% ~60% of the peak strength [28–30].


## 2.3. Characteristic Law of Rock Energy Evolution

It is assumed that the loading system is a closed system with no heat exchange with the external environment. In the loading system, external load input mechanical energy is absorbed by rock specimens, the rock specimens elastic deformation occur in the form of elastic strain energy stored within rock specimens. As the load goes on, the elastic strain energy stored by the rock specimens become less and less, and the energy absorbed at this time is mainly used to produce microcracks inside the rock specimens, namely, the part of energy is dissipated. In order to compare and analyze the law of energy evolution corresponding to different stages of stress-strain curves, axial strain is used as the horizontal axis, axial stress as the left axis, and energy density as the right axis. The energy evolution curves of the conventional triaxial compression test sandstone under different confining pressures are shown in Figure3.Figure 3
Energy evolution curve of sandstone under different confining pressure. (a) Confining pressure 5 MPa. (b) Confining pressure 10 MPa. (c) Confining pressure 15 MPa. (d) Confining pressure 20 MPa.
(a)(b)(c)(d)It can be seen that the energy evolution process can be divided into prepeak stage and postpeak stage according to the stress-strain curve in Figure3 and that different energy evolution curves under different confining pressures have similar characteristics.
(1)
As axial strain increases, total energyU increased exponentially before the axial stress peak and the growth rate decreased slightly after the stress peak. The growth rate is correlated with the properties of stress and strain. The maximum of U varied under different confining pressures. It increased with the increase of confining pressure(2)
Elastic strain energyUe indicates the releasable energy stored in the rock specimens, the evolution law is consistent with the stress-strain curve. In the prepeak stage, the difference between Ue and U was not great. After reaching the peak, Ue dropped rapidly. Also, the difference increased as the axial strain increased. The law of Ue is consistent with the stress-strain curve, which is positively associated with the confining pressure(3)
Dissipated energyUd indicates the energy consumed by the rock specimens during plastic deformation, crack expansion and energy release, which is irreversible. The evolution trend of Ud was slow growth before the peak but rapid growth after the peak. It grew almost exponentially and quickly became a major component of energy. The maximum of Ud varied under different confining pressures and increased with the increase in confining pressure. After the peak, the growth rate of Ud also increased slightly with the increase in confining pressureTo study the transformation of the energy in different stages, the corresponding energy characteristics of each stage are drawn in the bar graphs, and Figure4 shows the energy distribution characteristics of different stages.Figure 4
Energy distribution characteristics at different stages. (a) Prepeak stage. (b) Postpeak stage.
(a)(b)As can be seen in Figure4(a), the energy absorbed in the prepeak stage was mainly converted to elastic strain energy, accounting for about 72% ~80% of the total energy, while dissipated energy remained 18% to 20%. As confining pressure increased, the amount of the stored elastic strain energy became larger. The elastic strain energy was 464 kJ·m-3 when confining pressure was 5 MPa, and when the confining pressure increased to 20 MPa, the elastic strain energy was 1102.6 kJ·m-3, with an apparent confining pressure effect. At this stage, the elastic strain energy generated by rock specimen’s deformation can be continuously stored, while energy is consumed for the closure of cracks inside the rock specimens and the initiation and propagation of microcracks. As can be seen From Figure 4(b), the vast majority of the energy absorbed in the postpeak stage was transferred to dissipated energy, accounting for about 87% of the total energy. As the confining pressure increased, dissipated energy also increased. When confining pressure was 5 MPa, dissipated energy was 787 kJ·m-3. When confining pressure increased to 20 MPa, elastic strain energy was 1733 kJ·m-3. Elastic strain energy only accounted for a very small proportion. In this stage, a large number of macroscopic cracks form in the rock specimens and the total energy is mainly used for the formation and penetration of the macroscopic cracks. However, due to the existence of confining pressure, the rock specimens have certain residual strength, so a small amount of elastic strain energy can still be stored.


## 3. Energy Transformation of Roadway Surrounding Rock Numerical Simulation Method and Its Implementation

### 3.1. Energy Model of Roadway Surrounding Rock

Figure5 shows the mechanical model diagram of the roadway surrounding rock. Figure 5(a) is the schematic diagram of the roadway excavation, Q1 represents the underground space; Q2 represents roadway surrounding rock; and Q3 is roadway. The stress state of arbitrary point of surrounding rock is shown in Figure 5(b), which is represented by the stress component of the three orthogonal faces of the unit body. Ti indicates the stress component on arbitrary side. τij is the stress state of the surface, μj is the normal unit vector of the surface, and the relationships among the three satisfies Ti=μjτiji,j=1,2,3.Figure 5
Mechanical model of roadway surrounding rock. (a) Schematic diagram of roadway surrounding rock. (b) The stress state of any point of surrounding rock.
(a)(b)Supposing the stable state before the roadway excavation is State I, and the stable state of the roadway surrounding rock after the stress adjustment is State II. The energy transfornation process of roadway surrounding rock from State I to State II is shown in Figure6. Energy input consists of two parts. One part is the work done by underground space Q1 on surrounding rock Q2, which is called the external force work. The other part is the work done by physical strength in the unit body of surrounding rock Q2, which is called internal work. Energy accumulation consists of elastic strain energy, and the main source is the difference between the elastic strain energy before and after the excavation, namely, the difference between UI and UII. Energy dissipation consists of two parts. The first part is elastic strain energy Um stored in the excavated part of roadway surrounding rock Q2 and the second part is the plastic energy generated by the plastic deformation of surrounding rock.Figure 6
Schematic diagram of surrounding rock energy evolution. (a) Schematic diagram of energy input. (b) Schematic diagram of energy accumulation and dissipation.
(a)(b)


### 3.2. The Energy Model in FLAC3D

In FLAC3D the elastic strain energy and the dissipated plastic energy of the unit body containing mechanical model can be calculated. The whole process is solved incrementally: the motion equation of grid points and the stress strain of the region are solved in each time step. In the process where the system reaches equilibrium, the increment of the energy component can be determined by the stress-strain equation [31–33]. Elastic strain energy We of the unit body can be expressed as
(3)We=V2σijdσijd2G+σ¯K,where V is volume, σdij is partial stress, σ¯ is mean stress, and G and K are bulk modulus and shear modulus, respectively.Elastic strain energy incrementΔWe is the difference between the elastic energy in successive time steps, which can be divided into shear elastic energy increment ΔWes and volume elastic energy increment ΔWev. The expression is as follows:
(4)ΔWe=We′−We=ΔWes+ΔWev.Total shear strain energy incrementΔWTs and total volume strain energy increment ΔWTb can be obtained through average stress, respectively
(5)ΔWTs=V2σii+σii′eii+2σij+σij′eij,ΔWTv=3V2σ¯+σ¯′e¯.Plastic strain energy incrementΔWp can be divided into shear plastic energy increment ΔWps and volume plastic energy increment ΔWp, which is the difference between total strain energy and elastic strain energy. The expression is shown as follows:
(6)ΔWp=ΔWps+ΔWpv=ΔWTs−ΔWes+ΔWTv−ΔWev.The above energy formula is edited and transferred in the fish language built in FLAC3D so as to obtain such energy evolution characteristics of roadway surrounding rock as elastic strain energy and plastic strain energy.


### 3.3. Case of Study

Take the left fourth coal ventilation roadway in the 93rd coal of Xinjian coal mine as an example. A calculation model is established. In order to ensure the accuracy of the model, the stratigraphic parameters are appropriately simplified. The histogram of each stratum is shown in Figure7. The left fourth coal ventilation roadway is a rectangular roadway which is 3 m in width and 2 m in hight. The dimensions of the model areFigure 7
Histogram of coal strata.x×y×z=35m×20m×30m. The x-axis direction is trend, the y-axis direction is inclination, and the z-axis direction is gravity. There are 57576 nodes and 52400 unit bodies.The constitutive relationship of the model adopts Mohr-Coulomb criterion. Displacement constraints are made on the four sides and bottom surface of the model. According to the stress test report and other data,σz=10.34 MPa and σx=12.36 MPa are applied to the model. Model schematic and boundary conditions is shown in Figure 8. The mechanical parameters of surrounding rock are shown in Table 1.Figure 8
Model schematic and boundary conditions.Table 1
Mechanical parameters of surrounding rock.
Rock formationBulk modulus/GPaShear modulus/GPaCohesion/MPaInternal frictional angle/°Tensile strength/MPaDensity/kg·m-3Medium sandstone7.25.44.46323.092568Siltstone-fine sandstone interbedding8.56.33.54313.03264093# coal5.33.12.36281.681395Midfine sandstone10.37.83.25322.622608Medium sandstone9.57.33.89304.202568Fine sandstone9.86.24.02322.892614Figure9 shows the nephogram of the distribution of the rock plastic zone of roadway surrounding rock. The shear failure of the two ribs and the top and bottom corner of the roadway appeared in different degrees and distributed roughly symmetrically. Compared with top corners, the destructed range on bottom corners was larger. That is because the stress on the two ribs of roadway is transferred to the bottom corners, causing worse shear damage to bottom corners.Figure 9
Plastic distribution nephogram of surrounding rock.In the process of roadway excavation, the initial stress state of surrounding rock changes and the energy also changes accordingly. Figure10 is energy distribution nephogram of surrounding rock. As can be seen from Figure 10(a), the elastic strain energy of surrounding rock spreaded around the boundary of roadway. The spread range on the roof and floor was longer, which was roughly in the shape of strip with round ends. Both of ribs spreaded roughly in a circular way. The maximum of elastic strain energy density was 68.5 kJ·m-3, located in the position 1.2 m away from the two ribs of the roadway. As can be seen in Figure 10(b), the distribution range of plastic strain energy was consistent with plastic failure area, mainly distributing on the two ribs of the roadway, in the shape of semicircle, with a slightly larger bottom corners. Plastic strain energy spreaded from the both of ribs of the roadway. The spread range was about 1.2 m. The maximum of plastic strain energy density was 15.2 kJ·m-3, located in the boundary position of the two ribs. In fact, energy distribution of the roadway surrounding rock corresponds to plastic zone, which arises after surrounding rock reaches the yield state. This shows that, under those geological conditions, two ribs are the main sources of energy accumulation and dissipation.Figure 10
Energy density nephogram of roadway surrounding rock. (a) Elastic strain energy. (b) Plastic strain energy.
(a)(b)Figure11 shows energy characteristic change curve of the roadway ribs. The elastic strain energy density located within 1.2 m from the roadway surface was in the increase stage, rapidly increasing from 16.6 kJ·m-3 to 68.6 kJ·m-3. As the distance from the two ribs increased, elastic strain energy density decreased rapidly. When the distance increased to 6 m, the trend became stable gradually. Within the range 1.2 m from the surface of the roadway, plastic strain energy density decreased rapidly from 15.1 kJ·m-3 to about zero. That is because plastic deformation of the roadway arises on the two ribs. The range of plastic zone was roughly 1.2 m from the roadway surface. Plastic deformation of the roadway arose in this area, generating larger elastic strain energy and plastic strain energy. The deformation of surrounding rock outside the plastic zone decreased gradually and the elastic strain energy decreased gradually while the plastic strain energy was zero.Figure 11
Energy change curve of ribs of roadway.


## 3.1. Energy Model of Roadway Surrounding Rock

Figure5 shows the mechanical model diagram of the roadway surrounding rock. Figure 5(a) is the schematic diagram of the roadway excavation, Q1 represents the underground space; Q2 represents roadway surrounding rock; and Q3 is roadway. The stress state of arbitrary point of surrounding rock is shown in Figure 5(b), which is represented by the stress component of the three orthogonal faces of the unit body. Ti indicates the stress component on arbitrary side. τij is the stress state of the surface, μj is the normal unit vector of the surface, and the relationships among the three satisfies Ti=μjτiji,j=1,2,3.Figure 5
Mechanical model of roadway surrounding rock. (a) Schematic diagram of roadway surrounding rock. (b) The stress state of any point of surrounding rock.
(a)(b)Supposing the stable state before the roadway excavation is State I, and the stable state of the roadway surrounding rock after the stress adjustment is State II. The energy transfornation process of roadway surrounding rock from State I to State II is shown in Figure6. Energy input consists of two parts. One part is the work done by underground space Q1 on surrounding rock Q2, which is called the external force work. The other part is the work done by physical strength in the unit body of surrounding rock Q2, which is called internal work. Energy accumulation consists of elastic strain energy, and the main source is the difference between the elastic strain energy before and after the excavation, namely, the difference between UI and UII. Energy dissipation consists of two parts. The first part is elastic strain energy Um stored in the excavated part of roadway surrounding rock Q2 and the second part is the plastic energy generated by the plastic deformation of surrounding rock.Figure 6
Schematic diagram of surrounding rock energy evolution. (a) Schematic diagram of energy input. (b) Schematic diagram of energy accumulation and dissipation.
(a)(b)


## 3.2. The Energy Model in FLAC3D

In FLAC3D the elastic strain energy and the dissipated plastic energy of the unit body containing mechanical model can be calculated. The whole process is solved incrementally: the motion equation of grid points and the stress strain of the region are solved in each time step. In the process where the system reaches equilibrium, the increment of the energy component can be determined by the stress-strain equation [31–33]. Elastic strain energy We of the unit body can be expressed as
(3)We=V2σijdσijd2G+σ¯K,where V is volume, σdij is partial stress, σ¯ is mean stress, and G and K are bulk modulus and shear modulus, respectively.Elastic strain energy incrementΔWe is the difference between the elastic energy in successive time steps, which can be divided into shear elastic energy increment ΔWes and volume elastic energy increment ΔWev. The expression is as follows:
(4)ΔWe=We′−We=ΔWes+ΔWev.Total shear strain energy incrementΔWTs and total volume strain energy increment ΔWTb can be obtained through average stress, respectively
(5)ΔWTs=V2σii+σii′eii+2σij+σij′eij,ΔWTv=3V2σ¯+σ¯′e¯.Plastic strain energy incrementΔWp can be divided into shear plastic energy increment ΔWps and volume plastic energy increment ΔWp, which is the difference between total strain energy and elastic strain energy. The expression is shown as follows:
(6)ΔWp=ΔWps+ΔWpv=ΔWTs−ΔWes+ΔWTv−ΔWev.The above energy formula is edited and transferred in the fish language built in FLAC3D so as to obtain such energy evolution characteristics of roadway surrounding rock as elastic strain energy and plastic strain energy.


## 3.3. Case of Study

Take the left fourth coal ventilation roadway in the 93rd coal of Xinjian coal mine as an example. A calculation model is established. In order to ensure the accuracy of the model, the stratigraphic parameters are appropriately simplified. The histogram of each stratum is shown in Figure7. The left fourth coal ventilation roadway is a rectangular roadway which is 3 m in width and 2 m in hight. The dimensions of the model areFigure 7
Histogram of coal strata.x×y×z=35m×20m×30m. The x-axis direction is trend, the y-axis direction is inclination, and the z-axis direction is gravity. There are 57576 nodes and 52400 unit bodies.The constitutive relationship of the model adopts Mohr-Coulomb criterion. Displacement constraints are made on the four sides and bottom surface of the model. According to the stress test report and other data,σz=10.34 MPa and σx=12.36 MPa are applied to the model. Model schematic and boundary conditions is shown in Figure 8. The mechanical parameters of surrounding rock are shown in Table 1.Figure 8
Model schematic and boundary conditions.Table 1
Mechanical parameters of surrounding rock.
Rock formationBulk modulus/GPaShear modulus/GPaCohesion/MPaInternal frictional angle/°Tensile strength/MPaDensity/kg·m-3Medium sandstone7.25.44.46323.092568Siltstone-fine sandstone interbedding8.56.33.54313.03264093# coal5.33.12.36281.681395Midfine sandstone10.37.83.25322.622608Medium sandstone9.57.33.89304.202568Fine sandstone9.86.24.02322.892614Figure9 shows the nephogram of the distribution of the rock plastic zone of roadway surrounding rock. The shear failure of the two ribs and the top and bottom corner of the roadway appeared in different degrees and distributed roughly symmetrically. Compared with top corners, the destructed range on bottom corners was larger. That is because the stress on the two ribs of roadway is transferred to the bottom corners, causing worse shear damage to bottom corners.Figure 9
Plastic distribution nephogram of surrounding rock.In the process of roadway excavation, the initial stress state of surrounding rock changes and the energy also changes accordingly. Figure10 is energy distribution nephogram of surrounding rock. As can be seen from Figure 10(a), the elastic strain energy of surrounding rock spreaded around the boundary of roadway. The spread range on the roof and floor was longer, which was roughly in the shape of strip with round ends. Both of ribs spreaded roughly in a circular way. The maximum of elastic strain energy density was 68.5 kJ·m-3, located in the position 1.2 m away from the two ribs of the roadway. As can be seen in Figure 10(b), the distribution range of plastic strain energy was consistent with plastic failure area, mainly distributing on the two ribs of the roadway, in the shape of semicircle, with a slightly larger bottom corners. Plastic strain energy spreaded from the both of ribs of the roadway. The spread range was about 1.2 m. The maximum of plastic strain energy density was 15.2 kJ·m-3, located in the boundary position of the two ribs. In fact, energy distribution of the roadway surrounding rock corresponds to plastic zone, which arises after surrounding rock reaches the yield state. This shows that, under those geological conditions, two ribs are the main sources of energy accumulation and dissipation.Figure 10
Energy density nephogram of roadway surrounding rock. (a) Elastic strain energy. (b) Plastic strain energy.
(a)(b)Figure11 shows energy characteristic change curve of the roadway ribs. The elastic strain energy density located within 1.2 m from the roadway surface was in the increase stage, rapidly increasing from 16.6 kJ·m-3 to 68.6 kJ·m-3. As the distance from the two ribs increased, elastic strain energy density decreased rapidly. When the distance increased to 6 m, the trend became stable gradually. Within the range 1.2 m from the surface of the roadway, plastic strain energy density decreased rapidly from 15.1 kJ·m-3 to about zero. That is because plastic deformation of the roadway arises on the two ribs. The range of plastic zone was roughly 1.2 m from the roadway surface. Plastic deformation of the roadway arose in this area, generating larger elastic strain energy and plastic strain energy. The deformation of surrounding rock outside the plastic zone decreased gradually and the elastic strain energy decreased gradually while the plastic strain energy was zero.Figure 11
Energy change curve of ribs of roadway.


## 4. Roadway Surrounding Rock Control Technology Based on Energy Transformation

Roadway is in a stable state before excavation, while stress distribution of roadway is adjusted after excavation. Stress concentration may occur in some positions, where a lot of energy may accumulate. According to the energy balance law [34–36], the accumulated energy is released substantially; plastic failure may arise in surrounding rock. The energy in this process is irreversible. The internal energy conversion of the fully elastic rock mass can be expressed as
(7)Wc+Um=Uc+Wr,where Wc is the work done on rock mass by roadway excavation stress. Um is the strain energy generated from the excavated rock mass; Uc is the strain energy accumulated for roadway excavation; and Wc is the elastic energy released by roadway excavation.In the incompletely elastic roadway,Wr is approximate constant and Um and Uc are fixed values, which are related to the geological conditions of roadway surrounding rock. Therefore, the internal energy transformation of incompletely elastic rock mass can be expressed as
(8)Um=Uc+Wr+Wn+Wf,where Uc is the strain energy accumulated by the incompletely elastic roadway excavation, Wn is the energy released when the roadway is damaged and Wf is the energy absorbed by the support.Based on the energy characteristics and rock energy evolution process, the control technology principle of roadway surrounding rock stability is proposed, as is shown in Figure12. The first is reducing the elastic strain energy of surrounding rock. (1) By increasing the extension amount of the support. Part of the roadway energy is transferred to the support body, thus improving the energy release of surrounding rock; (2) By optimizing the layout of the roadway. Select the positions with good geological conditions for roadway construction, try to avoid the areas with such stress concentration as faults and geological structures to avoid stress concentration, thus reducing the energy accumulation of surrounding rock. The second is increasing the plastic strain energy of surrounding rock by setting weak structures, such as borehole pressure relief and other measures, to consume part of the energy in the coal body, thus improving the energy dissipation of surrounding rock.Figure 12
Roadway control technology principle based on the energy transformation.


## 5. Industrial Tests

### 5.1. Project Overview

The strike length of the right second working surface of the 91st coal of Xinjian coal mine is 800 m in depth, and the average dip length is 130 m. The average thickness of the coal seam is 1.4 m and the mean obliquity is 11°. Obliquity changes less and it tends to become larger from the east to the west. The occurrence of coal seam is stable and the coal type is 1/3 coking coal, which has stable structure and clear layer structure. The roof and floor is sandstone, which is gray and has high hardness.The cross section of the transport roadway is trapezoidal, 4.0 m wide and the central axis height of 3.2 m. Joint anchor bolt-cable support is adopted. The anchor cable is 5.0 m long, the diameter is 17.8 mm and the row spacing is2.0m×1.2m. The bolt is 2.4 m long, with a 20.0 mm diameter and 1.0m×1.2m row spacing. The section support diagram is shown in Figure 13.Figure 13
Roadway section support map.


### 5.2. Stability Control Scheme of Roadway Surrounding Rock

In the actual roadway engineering, the optimization of roadway layout is greatly limited by the geological conditions and excavation influence, which is only applicable to the initial stage of roadway construction. It is difficult to set weak structures for construction, which has high requirements for the size, height, and angle of weak structures. The energy release in roadway surrounding rock is mostly released in the form of deformation energy. For the roadway with specific geological conditions, the total amount of released energy is certain. Energy can be released on the support to reduce the energy amount on surrounding rock. Therefore, roadway stability control is conducted mainly in terms of improving the energy release of surrounding rock.


#### 5.2.1. Roadway Support Measures

(1) Control roof stability with high-density long anchor cable. On the basis of the original support, adjust the parameters of the support, strengthen the supporting role of the anchor cable, and improve the integrity of surrounding rock so that the surrounding rock and the support body can form a complete bearing structure. Adopt high density long anchor cable, 7300 mm long, with a diameter of 21.8 mm, whose row spacing is1000mm×1000mm, arranged alternatively with bolts. (2) Replace the ribs bolt with the short anchor cable with a large diameter. Due to the limitation of length, the control range to the two ribs of the ordinary bolt is limited. The ordinary short bolt is replaced with the short anchor cable with a large diameter. Its length is greater than the ordinary short bolt, which can increase the control range of the surrounding rock on the ribs. The pretightening force of the large-diameter anchor cable is higher than the ordinary anchor cable, which can effectively control the loosening and destruction of the surrounding rock on the ribs. The short anchor cable is 3000 mm long and its diameter is 21.8 mm and row spacing is 1000mm×1000mmAccording to the above support measures, the support parameters are designed. The support map of the optimized roadway section is shown in Figure14.Figure 14
Optimized roadway section support map.


#### 5.2.2. Analysis of Support Optimization

Take the ventilation roadway in the 91st coal of Xinjian Coal Mine as the engineering background. According to the data of the mine, a calculation model is established. In order to ensure the accuracy of the model, the strati parameters are appropriately simplified. The roadway is 4.0 m wide and the central axis is 3.2 m high. The dimensions of the model arex×y×z=40m×20m×30m. The constitutive model adopts the Mohr-Coulomb criterion. Displacement constrains are conducted on the four sides and the bottom of the model. According to the geological data, in suit stress σz=10.34MPa, σx=11.63 MPa, σy=7.15MPa is applied to the model. The mechanical parameters of the surrounding rock are shown in Table 2.Table 2
Mechanical parameters and thickness of surrounding rock.
Rock formationBulk modulus/GPaShear modulus/GPaCohesion/MPaInternal frictional angle/°Tensile strength/MPaDensity/kg·m-3Thickness/mFine sandstone417.210.44.46322.1825682.35Siltite218.511.33.54312.3526401.09Middle sandstone319.812.12.36281.9813954.55Fine sandstone310.39.83.25322.6226083.2Siltite119.511.33.89303.2025681.3891Coal7.43.52.38251.6813981.35Middle sandstone 219.812.12.36281.9813951.56Fine sandstone 210.59.53.25322.6226082.290 coal7.33.42.19261.5213960.85Middle sandstone 119.812.12.36281.9813951.0Fine sandstone 110.29.33.25322.6226085.95Figure15 shows the energy nephograms before and after the support optimization. The energy distribution of the roof and floor of roadway was relatively uniform, and the energy distribution range of the roof was slightly greater than floor. The high energy area connection of the roadway ribs was roughly the same as the seam inclination, and the high energy area on the right rib was larger. Before and after the support optimization, the maximum of elastic strain energy decreased significantly and the maximum of elastic strain energy of the two ribs decreased from 193.3 kJ·m-3 to 159.0 kJ·m-3, and the maximum of elastic strain energy on the roof and floor decreased from 114.2 kJ·m-3 to 95.3 kJ·m-3, decreasing 17.7% and 16.5%, respectively. After the support optimization, the extension amount of the support increased, and the energy of roadway surrounding rock can be transferred to the support body. Thus, the energy release of roadway surrounding rock improves and the elastic strain energy is also reduced.Figure 15
Elastic strain energy nephogram before and after optimization. (a) Original support. (b) After optimization.
(a)(b)


### 5.3. Validation of Effect

In order to verify the rationality of the supporting parameters, monitoring points were selected on the right rib, roof, and floor of the roadway to monitor the deformation amount. They were monitored every 3 to 5 days, which lasted 60 days. The displacement change curve of the surface before and after the roadway was optimized is shown in Figure16. The surface displacement amount of the roadway before and after the optimization dropped apparently. Before the optimization, the displacement amount of the roadway surface showed a growth trend. The floor had an obvious slowing trend while the roof and the right rib still had a growth trend. The maximum of roof subsidence, floor heave and shrinkage of the right rib were 247 mm, 120 mm, and 292 mm, respectively. After the optimization, the increase rate of roadway surface displacement became slow on the whole, and the effect of floor was particularly obvious, which gradually became stable after 45 days. Each displacement amount decreased apparently: roof subsidence decreased to 62 mm, floor heave dropped to 26 mm, and right rib shrinkage dropped to 94 mm.Figure 16
Surface displacement curve of roadway. (a) Before the optimization. (b) After the optimization.
(a)(b)


## 5.1. Project Overview

The strike length of the right second working surface of the 91st coal of Xinjian coal mine is 800 m in depth, and the average dip length is 130 m. The average thickness of the coal seam is 1.4 m and the mean obliquity is 11°. Obliquity changes less and it tends to become larger from the east to the west. The occurrence of coal seam is stable and the coal type is 1/3 coking coal, which has stable structure and clear layer structure. The roof and floor is sandstone, which is gray and has high hardness.The cross section of the transport roadway is trapezoidal, 4.0 m wide and the central axis height of 3.2 m. Joint anchor bolt-cable support is adopted. The anchor cable is 5.0 m long, the diameter is 17.8 mm and the row spacing is2.0m×1.2m. The bolt is 2.4 m long, with a 20.0 mm diameter and 1.0m×1.2m row spacing. The section support diagram is shown in Figure 13.Figure 13
Roadway section support map.


## 5.2. Stability Control Scheme of Roadway Surrounding Rock

In the actual roadway engineering, the optimization of roadway layout is greatly limited by the geological conditions and excavation influence, which is only applicable to the initial stage of roadway construction. It is difficult to set weak structures for construction, which has high requirements for the size, height, and angle of weak structures. The energy release in roadway surrounding rock is mostly released in the form of deformation energy. For the roadway with specific geological conditions, the total amount of released energy is certain. Energy can be released on the support to reduce the energy amount on surrounding rock. Therefore, roadway stability control is conducted mainly in terms of improving the energy release of surrounding rock.


### 5.2.1. Roadway Support Measures

(1) Control roof stability with high-density long anchor cable. On the basis of the original support, adjust the parameters of the support, strengthen the supporting role of the anchor cable, and improve the integrity of surrounding rock so that the surrounding rock and the support body can form a complete bearing structure. Adopt high density long anchor cable, 7300 mm long, with a diameter of 21.8 mm, whose row spacing is1000mm×1000mm, arranged alternatively with bolts. (2) Replace the ribs bolt with the short anchor cable with a large diameter. Due to the limitation of length, the control range to the two ribs of the ordinary bolt is limited. The ordinary short bolt is replaced with the short anchor cable with a large diameter. Its length is greater than the ordinary short bolt, which can increase the control range of the surrounding rock on the ribs. The pretightening force of the large-diameter anchor cable is higher than the ordinary anchor cable, which can effectively control the loosening and destruction of the surrounding rock on the ribs. The short anchor cable is 3000 mm long and its diameter is 21.8 mm and row spacing is 1000mm×1000mmAccording to the above support measures, the support parameters are designed. The support map of the optimized roadway section is shown in Figure14.Figure 14
Optimized roadway section support map.


### 5.2.2. Analysis of Support Optimization

Take the ventilation roadway in the 91st coal of Xinjian Coal Mine as the engineering background. According to the data of the mine, a calculation model is established. In order to ensure the accuracy of the model, the strati parameters are appropriately simplified. The roadway is 4.0 m wide and the central axis is 3.2 m high. The dimensions of the model arex×y×z=40m×20m×30m. The constitutive model adopts the Mohr-Coulomb criterion. Displacement constrains are conducted on the four sides and the bottom of the model. According to the geological data, in suit stress σz=10.34MPa, σx=11.63 MPa, σy=7.15MPa is applied to the model. The mechanical parameters of the surrounding rock are shown in Table 2.Table 2
Mechanical parameters and thickness of surrounding rock.
Rock formationBulk modulus/GPaShear modulus/GPaCohesion/MPaInternal frictional angle/°Tensile strength/MPaDensity/kg·m-3Thickness/mFine sandstone417.210.44.46322.1825682.35Siltite218.511.33.54312.3526401.09Middle sandstone319.812.12.36281.9813954.55Fine sandstone310.39.83.25322.6226083.2Siltite119.511.33.89303.2025681.3891Coal7.43.52.38251.6813981.35Middle sandstone 219.812.12.36281.9813951.56Fine sandstone 210.59.53.25322.6226082.290 coal7.33.42.19261.5213960.85Middle sandstone 119.812.12.36281.9813951.0Fine sandstone 110.29.33.25322.6226085.95Figure15 shows the energy nephograms before and after the support optimization. The energy distribution of the roof and floor of roadway was relatively uniform, and the energy distribution range of the roof was slightly greater than floor. The high energy area connection of the roadway ribs was roughly the same as the seam inclination, and the high energy area on the right rib was larger. Before and after the support optimization, the maximum of elastic strain energy decreased significantly and the maximum of elastic strain energy of the two ribs decreased from 193.3 kJ·m-3 to 159.0 kJ·m-3, and the maximum of elastic strain energy on the roof and floor decreased from 114.2 kJ·m-3 to 95.3 kJ·m-3, decreasing 17.7% and 16.5%, respectively. After the support optimization, the extension amount of the support increased, and the energy of roadway surrounding rock can be transferred to the support body. Thus, the energy release of roadway surrounding rock improves and the elastic strain energy is also reduced.Figure 15
Elastic strain energy nephogram before and after optimization. (a) Original support. (b) After optimization.
(a)(b)


## 5.2.1. Roadway Support Measures

(1) Control roof stability with high-density long anchor cable. On the basis of the original support, adjust the parameters of the support, strengthen the supporting role of the anchor cable, and improve the integrity of surrounding rock so that the surrounding rock and the support body can form a complete bearing structure. Adopt high density long anchor cable, 7300 mm long, with a diameter of 21.8 mm, whose row spacing is1000mm×1000mm, arranged alternatively with bolts. (2) Replace the ribs bolt with the short anchor cable with a large diameter. Due to the limitation of length, the control range to the two ribs of the ordinary bolt is limited. The ordinary short bolt is replaced with the short anchor cable with a large diameter. Its length is greater than the ordinary short bolt, which can increase the control range of the surrounding rock on the ribs. The pretightening force of the large-diameter anchor cable is higher than the ordinary anchor cable, which can effectively control the loosening and destruction of the surrounding rock on the ribs. The short anchor cable is 3000 mm long and its diameter is 21.8 mm and row spacing is 1000mm×1000mmAccording to the above support measures, the support parameters are designed. The support map of the optimized roadway section is shown in Figure14.Figure 14
Optimized roadway section support map.


## 5.2.2. Analysis of Support Optimization

Take the ventilation roadway in the 91st coal of Xinjian Coal Mine as the engineering background. According to the data of the mine, a calculation model is established. In order to ensure the accuracy of the model, the strati parameters are appropriately simplified. The roadway is 4.0 m wide and the central axis is 3.2 m high. The dimensions of the model arex×y×z=40m×20m×30m. The constitutive model adopts the Mohr-Coulomb criterion. Displacement constrains are conducted on the four sides and the bottom of the model. According to the geological data, in suit stress σz=10.34MPa, σx=11.63 MPa, σy=7.15MPa is applied to the model. The mechanical parameters of the surrounding rock are shown in Table 2.Table 2
Mechanical parameters and thickness of surrounding rock.
Rock formationBulk modulus/GPaShear modulus/GPaCohesion/MPaInternal frictional angle/°Tensile strength/MPaDensity/kg·m-3Thickness/mFine sandstone417.210.44.46322.1825682.35Siltite218.511.33.54312.3526401.09Middle sandstone319.812.12.36281.9813954.55Fine sandstone310.39.83.25322.6226083.2Siltite119.511.33.89303.2025681.3891Coal7.43.52.38251.6813981.35Middle sandstone 219.812.12.36281.9813951.56Fine sandstone 210.59.53.25322.6226082.290 coal7.33.42.19261.5213960.85Middle sandstone 119.812.12.36281.9813951.0Fine sandstone 110.29.33.25322.6226085.95Figure15 shows the energy nephograms before and after the support optimization. The energy distribution of the roof and floor of roadway was relatively uniform, and the energy distribution range of the roof was slightly greater than floor. The high energy area connection of the roadway ribs was roughly the same as the seam inclination, and the high energy area on the right rib was larger. Before and after the support optimization, the maximum of elastic strain energy decreased significantly and the maximum of elastic strain energy of the two ribs decreased from 193.3 kJ·m-3 to 159.0 kJ·m-3, and the maximum of elastic strain energy on the roof and floor decreased from 114.2 kJ·m-3 to 95.3 kJ·m-3, decreasing 17.7% and 16.5%, respectively. After the support optimization, the extension amount of the support increased, and the energy of roadway surrounding rock can be transferred to the support body. Thus, the energy release of roadway surrounding rock improves and the elastic strain energy is also reduced.Figure 15
Elastic strain energy nephogram before and after optimization. (a) Original support. (b) After optimization.
(a)(b)


## 5.3. Validation of Effect

In order to verify the rationality of the supporting parameters, monitoring points were selected on the right rib, roof, and floor of the roadway to monitor the deformation amount. They were monitored every 3 to 5 days, which lasted 60 days. The displacement change curve of the surface before and after the roadway was optimized is shown in Figure16. The surface displacement amount of the roadway before and after the optimization dropped apparently. Before the optimization, the displacement amount of the roadway surface showed a growth trend. The floor had an obvious slowing trend while the roof and the right rib still had a growth trend. The maximum of roof subsidence, floor heave and shrinkage of the right rib were 247 mm, 120 mm, and 292 mm, respectively. After the optimization, the increase rate of roadway surface displacement became slow on the whole, and the effect of floor was particularly obvious, which gradually became stable after 45 days. Each displacement amount decreased apparently: roof subsidence decreased to 62 mm, floor heave dropped to 26 mm, and right rib shrinkage dropped to 94 mm.Figure 16
Surface displacement curve of roadway. (a) Before the optimization. (b) After the optimization.
(a)(b)


## 6. Discussion

In this paper, based on elastic strain energy and plastic strain energy, the evolution law of sandstone energy characteristics and roadway surrounding rock energy evolution law were obtained, respectively, through laboratory test and numerical simulation, and the stability control technology principle of roadway surrounding rock based on energy transformation is proposed. According to the geological conditions, the stability control of roadway surrounding rock was carried out. It shows that the technique principle is suitable for surrounding rock stability control, and the technique of surrounding rock stability control is a supplement to mechanics from the perspective of energy, and has a good application prospect.In the process of rock deformation and failure under external force, the energy changes all the time. In this paper, the energy evolution law of rock specimens in the process of failure is studied by elastic strain energy and dissipation energy, ignoring energy characteristic parameters such as frictional dissipation energy used between cracks in the process of energy evolution, thus there are some errors in the description of energy evolution in the whole process of deformation and failure.The stability of roadway surrounding rock can be reflected by the energy transformation of surrounding rock. The numerical model of roadway surrounding rock energy established in this paper is based on the elastic strain energy and plastic strain energy of surrounding rock, which has good application effect, but lacks the verification of field application. Therefore, the energy of roadway surrounding rock can be quantified by combining seismological principle.From the perspective of energy transformation, the principle of roadway surrounding rock stability control technology is put forward. Due to the limitation of geological conditions, this paper only carried out field construction aiming at improving the energy release principle of surrounding rock, and achieved good application effect. The effect of improving the energy dissipation of surrounding rock, reducing the energy accumulation, and the synergistic effect of each other have not been verified in the field, so corresponding research should be carried out.


## 7. Conclusion

In this paper, based on the research of the energy evolution regularity of sandstone samples, the surrounding rock energy numerical model was established, and then the principle of roadway surrounding rock stability control technology is put forward. Field industrial test results show that the control effect is good. The main conclusions are as follows:(1)
In the process of rock deformation and failure, the total energy increased exponentially before the axial stress peak, mainly transferred to elastic strain energy, accounting for about 72% ~80% of the total energy and the dissipated energy remained 18% ~20%. The total energy growth rate decreased slightly in the postpeak stage. The vast majority of energy was transferred to dissipated energy, accounting for about 87% of the total energy(2)
Combined with energy transformation, put forward the principle of roadway surrounding rock stability control technology: one is to reduce the elastic strain energy of surrounding rock. On the one hand, by increasing the extension amount of support body, part of roadway energy is transferred to support body, thus improving the energy release of surrounding rock. On the other hand, by optimizing roadway layout, stress concentration is avoided and energy accumulation of surrounding rock is reduced. The second is to increase the plastic strain energy of surrounding rock, that is, to increase the energy dissipation of surrounding rock by setting weak structure(3)
After the optimization of support measures, the roadway surface displacement and its increase rate decreased significantly, roof subsidence decreased from 247 mm to 62 mm, floor heave decreased from 120 mm to 26 mm, and the right rib shrinkage decreased from 292 mm to 94 mm, that is, the floor heave control effect was particularly obvious. The deformation of surrounding rock gradually stabilized after 45 days



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*Source: 8938807-2023-01-19.xml*