Roadway deformation during riding mining in soft rock

International Journal of Mining Science and Technology 22 (2012) 539–544

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International Journal of Mining Science and Technology journal homepage: www.elsevier.com/locate/ijmst

Roadway deformation during riding mining in soft rock Zhao Guozhen a,b,⇑, Ma Zhanguo a,b, Zhu Qinghua c, Mao Xianbiao a,b, Feng Meimei a,b a

State Key Laboratory of Geomechanics and Deep Underground Engineering, China University of Mining & Technology, Xuzhou 221008, China School of Mechanics and Civil Engineering, China University of Mining & Technology, Xuzhou 221008, China c Zhangshuanglou Mine of Xuzhou Mine Group, Xuzhou 221616, China b

a r t i c l e

i n f o

Article history: Received 24 November 2011 Received in revised form 21 December 2011 Accepted 25 January 2012 Available online 7 July 2012 Keywords: Soft rock roadway Deformation mechanism Partition broken rock Viscoelastic-plastic analysis

a b s t r a c t ‘‘Riding mining’’ is a form of mining where the working face is located above the roadway and advances parallel to it. Riding mining in deep soft rock creates a particular set of problems in the roadway that include high stresses, large deformations, and support difficulties. Herein we describe a study of the rock deformation mechanism of a roadway as observed during riding mining in deep soft rock. Theoretical analysis, numerical simulations, and on site monitoring were used to examine this problem. The stress in the rock and the visco-elastic behavior of the rock are considered. Real time data, recorded over a period of 240 days, were taken from a 750 transportation roadway. Stress distributions in the rock surrounding the roadway were studied by comparing simulations to observations from the mine. The rock stress shows dynamic behavior as the working face advances. The pressure increases and then drops after peaking as the face advances. Both elastic and plastic deformation of the surrounding rock occurs. Plastic deformation provides a mechanism by which stress in the rock relaxes due to material flow. A way to rehabilitate the roadway is suggested that will help ensure mine safety. Ó 2012 Published by Elsevier B.V. on behalf of China University of Mining & Technology.

1. Introduction Gradual increases in mining depth have caused overall stress in the rock around the mined space to increase. Some developed systems in deep mining districts are now affected by dynamic pressure, in our country. With plenty of faults, easily expanded roof strata show numerous faults when construction meets water or other factors. Roadways surrounded by soft rock show the typical deformation characteristics of high stress, large deformations, and support difficulties [1–4]. So called ‘‘riding mining’’ where the working face is located over the roadway and mining proceeds in the same direction as the roadway has its own difficulties. The special structural features create dynamic pressure and rock stresses that result in displacements caused by these pressures, which frequently are severe. The original state of equilibrium is broken and this further increases the difficulty of supporting the structures [5]. If early, appropriate measures are not taken to maintain the structure the severe deformation in the surrounding rocks eventually leads to unstable failure of the roadway. This will then destroy the important mine transport systems, ventilation systems, pedestrian systems, and piping system. This poses a serious threat to normal mine production and to the safety of personnel. In addition, signif-

⇑ Corresponding author. Tel.: +86 13585478539. E-mail address: [email protected] (G. Zhao).

icant renovations of the roadway upset production schedules in the mine and also cause duplication of investment when previously installed structure must be re-constructed. This increases the pressure on costs and benefits. Therefore, research on roadway deformation in deep, soft rock related to riding mining is particularly important. This has a profound significance to a large number of mines that need effective ways and means of support. For them this work is important to achieve sustainable economic and efficient coal mine development [6,7].

2. Geological conditions In some mines from 70% to 80% of the roadway will be affected by mining. Cross mining and riding mining are two important techniques where dynamic pressure can affect the roadway (Fig. 1) [8]. A large number of key roadways have been damaged by the dynamic pressure from riding mining in a mine located in Xuzhou. There, the roadway was developed into a deep area [9]. The 750 roadway, now affected by riding mining of the fully mechanized caving face number 22107, has the most complicated conditions. The rock lithology is unfavorable for structural stability and requires the most demanding deformation control in the roadway. The 750 roadway is a key roadway for transport of coal out of this coal mine. It is nearly 800 m long. There is a belt conveyor with two turntable devices (11° and 14°) along it. One of the turntable devices is

2095-2686/$ - see front matter Ó 2012 Published by Elsevier B.V. on behalf of China University of Mining & Technology. http://dx.doi.org/10.1016/j.ijmst.2012.01.017

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G. Zhao et al. / International Journal of Mining Science and Technology 22 (2012) 539–544

Fig. 1. Simple diagrams showing cross- and riding-mining.

deployed in a triangular bifurcated passage. There are many faults and a water rich sandstone roof in the rock surrounding the roadway. The roof and floor are sand-shale, which expands easily and softens in direct contact with water. The minimum vertical distance between the mining face and the 750 roadway is 28.5 m and the minimum horizontal distance between one side of the mining face and the coal wall of the 750 roadway is about 18.0 m. Early on, the roadway used ordinary supports in the form of steel anchor mesh and shotcrete. However, after significant mining the roadway began to exhibit stress damage and huge deformations. So it is urgently in need of repair now [10,11].

3. Stress and deformation in the surrounding rock: visco-elastic analysis Rocks will show strong rheological properties in the long term under high stress. Hence the visco-elastic theory is an appropriate calculation model for the study of the rock surrounding the roadway. Rock stability during riding mining and rock failure under dynamic pressure can be understood through the use of this model. A modified Fenner formula indicates that the interface in the plastic area is always at a constant stress level. The interface is not pertinent to the matter of the radius of the plastic zone nor to the roadway support resistance [12–16]. This applies between the elastic and the plastic interface zone. The radius, R0, of the plastic zone is a function of time, t, which is to say that R0 changes over time. Consider the radius of a point at r = aR0(t). In the viscoelastic zone the point changes as a function of R0(t) and the stress state does not change over time. The calculation model is shown in Fig. 2.

The visco-elastic analysis of the roadway proceeds through the use of an analytical formula describing the visco-elastic zone and plastic deformation in the surrounding rock. The visco-elastic displacement and stress formulas are:

ur ¼ MR40aðtÞ

h

1 G1

i h i t t t t MR2 ðtÞ ð1  egret Þ þ G10 egret ¼ 4r0 G11 ð1  egret Þ þ G10 egret 2

rr ¼ P  M2 a12 ¼ P  M2 R0r2ðtÞ 2

rh ¼ P þ M2 a12 ¼ P þ M2 R0r2ðtÞ : ð1Þ As shown in Eq. (1) at the points where r changes (as a function of R0(t)) the stress does not change over time. But the displacement does change over time. The displacement in the plastic zone is given by:

up ðtÞ ¼

  t R20 ðtÞ M M g t ð1  egret Þ þ e ret : r 4G1 4G0

ð2Þ

The changing relationship between the plastic zone radius and time can be derived from Eq. (2):

2 R0 ðtÞ ¼ 4

312

upr0 ðtÞr 0

M 4G1

t

t

M gret ð1  egret Þ þ 4G e 0

5:

ð3Þ

The radius of the plastic zone, R0(t), and the support resistance, P0i ðtÞ, needed to maintain the plastic zone limit equilibrium can be derived from the modified Fenner formula:

8 sin /  1sin > / > 0 Mar0 > > P ðtÞ ¼ ðP þ Cctg/Þð1  sin /Þ  Cctg/ p > 4ur ðtÞ < i 0 " #12 > p > ur ðtÞr0 > 0 > ðtÞ ¼ R > t t 0 : M ð1egret Þþ M egret 4G1

ð4Þ

4G0

or

8 2 < Pi ðtÞ ¼ K c ½upr ðtÞ  u0  ¼ K c ½MbR0 ðtÞ  u0  0 4r 0 : Fig. 2. Model of the surrounding rock used for visco-elastic analysis (shading represents the plastic zone).

upr0 ðtÞ ¼

MbR20 ðtÞ 4r 0

ð5Þ

where Kc is the stiffness coefficient; Gc the shear modulus, and lc the Poisson’s ratio.

G. Zhao et al. / International Journal of Mining Science and Technology 22 (2012) 539–544

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Fig. 3. Damage of the roadway under early forms of support.

4. Surface deformation of the roadway

4.2. Roadway surface displacement analysis

4.1. Surface deformation

The deformation of the 750 transport roadway required the use the time sharing, phased pre-stressed grouting anchor techniques for roadway repair and reinforcement. The effects of the repairs were examined by setting up eight stations within the roadway around the mining face. Stations 1# and 2# were located in the north roadway. The 1# station was located about 40 m in front of the cut-hole of the fully mechanized mining face, number 22107. The distance between stations 2# and 1# is 100 m. Station 3# was located at the intersection of the roadway to analyze the impact of mining induced pressure at the intersection. Stations 4# to 8# were the auxiliary stations of the roadway intended to analyze mine pressure at different distances from the roadway to the fully-mechanized face. The rock pressures were monitored over a period of 8 months. The roadway surface displacement versus time data are shown in Fig. 4. This data show that:

The damage to the 750 roadway from geological factors and the dynamic pressure was more obvious over the early support period. The dynamic pressure is the main factor. Of course, there are many other additional factors, such as faults around the area, the lower intensity of the rock, and gravity acting on the deep rock itself. Some damage photos are given below. As shown in Fig. 3 it is clear that: (1) The roof of the roadway has suffered extensive damage and localized failure was particularly serious. Coal buried deeply at great pressure creates pressures that exceed the surrounding rock strength, resulting in tunnel collapse. (2) The coal wall of the roadway was damaged seriously. The main appearance is one of compression deformation and bending convergence. The 750 roadway is obviously affected by the dynamic pressure of mining. The pressure on the roof shifts into the surrounding rock making the pressure in this surrounding rock increase. The pressure squeezes the surrounding rocks and causes plastic deformation. A plastic wedge forms and slips inward to destroy the surrounding rock. The coal wall of the roadway is in a special geological formation where the expansion of soft rock by tectonic stress and water erosion has occurred. The lower strength of these materials allows the coal wall to undergo a great amount of drum deformation. (1) There are serious peeling sites in multiple locations of this roadway. The shrinkage of the concrete itself has created a spray layer that is in tension. When the normal tensile stress of the roadway exceeds the tensile strength of the concrete the spray layer fractures and spalls off, forming a bare spot in the rock.

(1) The deformation in the rock surrounding the transport roadway is almost zero in the early stages of monitoring. The impact from the No. 22107 face is very small at first. As the face moves forward the rock deformation also increases. Rock deformation is occurs for about 220 days after the face moves forward. Therefore, rheological behavior of the rock is the main factor involved with roadway deformation resulting from the influence of mining. (2) The deformation at station 3# was larger than at the other stations in the transportation roadway because this station is at the crossing of three roadways. This allowed the maximum affect from mining, and the stress concentration there, to be more significant. At station 3# the convergence of the roadway sides was 165 mm and the roof to floor convergence of the roadway was 174 mm. The roof to floor convergence was dominated by floor heave, which accounted for 70%, and the side nearest the face accounted for 65% of the side to side convergence.

Fig. 4. Surface displacement versus time.

G. Zhao et al. / International Journal of Mining Science and Technology 22 (2012) 539–544

Pressure (MPa)

Side of roof Lower part of sidewall

Center of roof Upper part of sidewall

2.5 2.0 1.5 1.0 0.5 0

1

7

13

19

25

31

37

43

Time (day)

(a) Measured anchor forces at station 6

2.5

Pressure (MPa)

542

Upper part of sidewall

Lower part of sidewall

7

31

2.0 1.5 1.0 0.5 0

1

13

19

25

37

43

Time (day)

(b) Measured anchor forces at station 7

Fig. 5. Forces in the anchors and anchor cables.

Fig. 6. Displacement versus time in the rock surrounding the deep roadway.

4.3. Monitoring the anchoring force Fig. 5 shows the data from stations 6# and 7#. As shown there: (1) The pressure in the rock surrounding the transport roadway fluctuates from the impact of the No.22107 face. Abrupt large, or small, changes were observed and some bolt anchors failed after a huge stress wave. (2) The pressure of the central roof on the side facing the coal fluctuated at station 6#, giving a peak value of 3.2 MPa. However, the pressure on the side away from the coal side fluctuated about 1 MPa and gradually stabilized. (3) The dynamic pressure affected the anchoring force of the anchor cable more at the beginning of the observation period. The initial anchoring force was over 1 MPa. The anchoring force in the anchor cable shows different degrees fluctuation over the initial 10 days. As the face moved forward the anchoring force increased or decreased and eventually stabilized after 20 days. The average force was 1.4 MPa. 5. Internal deformation of the rock 5.1. Internal displacement of the rock Internal deformation of the surrounding rock was examined by monitoring with a multi-point displacement method and a separation indicator. The displacement-time curves at locations of 1, 2, 3, and 4 m into the rock are shown in Fig. 6. Fig. 6 shows that: (1) This deep rock has a large displacement after being affected by dynamic pressure. The separation is similar both in the deep and shallow locations. The deeper separation was not more severe than the shallow one. The maximum deformation in the rock occurs at a depth of 1 m and the largest

convergence is 84.9 mm. As the depth increases further the deformation decreases, to only 55.2 mm at the 4 m depth. The surrounding rock moves significantly when the dynamic pressure causes flow deformation. However, the displacement of the surround rock decreases as the depth into it increases. (2) Comparing the displacement at stations 1# and 3# shows that station 3# had a slightly larger displacement. In addition, the deep displacement increased rapidly early in the observation period but then gradually stabilized at station three. Short term stationary periods are then sometimes followed by a period of change. This process shows that the surrounding activities are intermittent. 5.2. Rock internal stress The supports at the coal face pass stresses to the underlying coal layer when the coal face has advanced 20 m. A relief zone, a supporting stress zone, and a primary rock stress zone then form in the bottom coal layer. This has an impact on the stress and deformation of the lower roadway. Pressure versus time curves are shown in Fig. 7 for 3, 5, 7, and 11 m into this area. It is apparent that the measured, initial stresses were 2.0, 1.7, 3.8 and 4.0 MPa on the first day. As the face moved forward the pressure was re-tested and it obviously decreases and becomes stable at from 1.0 to 1.5 MPa over the next 5–14 days. After 15 days the stress shows small fluctuations at a position of 3 and 7 m, while the stress has become zero at 5 and 11 m. So, this stress does have some relief process after the face moves forward. 5.3. An analysis of rupture evolution The detection and analysis of the roadway deformation shows a regional fracture phenomenon distributed over intervals, see Fig. 8. This is very similar to the zonal disintegration phenomenon in rock [17]. The zonal disintegration phenomenon means that the rock surrounding the roadway develops alternating ruptured and

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3m 5m 7m 11 m

3 2

1.6

Pressure (MPa)

Pressure (MPa)

4

1 0

1

3

5

7

9

11 13 15 17 19 21 23 25 27

Time (day)

(a) Measured pressures at different depths

1.2 0.8

No.8 No.9 No.10 No.11

0.4 03

5

7

No.12 No.13 No.14 No.15 9

11

Drilling depth (m) (b) Measured pressures at different stations : day one

Fig. 7. Stresses in the underlying coal.

Fig. 8. Regional breakdown of the roadway.

other observation stations along the transportation roadway. The convergence of the roadway sides was 165 mm and the roof-to-floor convergence was 174 mm. (2) Deep rock has a large displacement after being affected by dynamic pressure. The separation shows a similar trend between deep and shallow points, and the deep separation is not more severe than the shallow separation. The displacement of the surrounding rock decreases as the depth increases. (3) The pressure in rock surrounding a transport roadway fluctuates under the influence of the face moving forward. Irregular large or small changes are seen and some bolt anchors see enough force to fail after a huge wave. (4) Deep stress has a relief process after it peaks when the face moves forward. The deep surrounding rock shows the zonal disintegration phenomenon. Acknowledgments This study is Supported by the National Natural Science Foundation of China (Nos. 50834005 and 51074163), the Ministry of Education Support Program for New Century Excellent of China (No. NCET-08-0837), and the Fundamental Research Funds for the Central Universities of China, and Youth Science and Technology Foundation of China University of Mining and Technology (No. 2010QNB25). We are grateful to our peer reviewers and the journal editors. References

Fig. 9. Maximum principal stress r1 contours in surrounding rock (50 days).

non-ruptured zones on the roadway sides and the front by the face. This has been seen before when a chamber, or roadway, is excavated in deep underground rock. It is associated with structural level failure in the rock. Since a deeply buried roadway is surrounded by highly stressed rock this extensive but regular behavior reveals another balancing process that allows a new stable state to form in the deep rock [18]. A numerical model using FLAC software code was used to model the surrounding rock [19–23]. The parameters are: a roadway depth, H, of 750 m, a vertical distance, h, of 20 m, and a horizontal distance, b, of 18 m. The maximum principal stress contours, r1, in a roadway constructed in soft rock during riding mining at different time periods is shown here as Fig. 9. 6. Conclusions (1) Rheological behavior is the main factor involved in roadway deformation. The deformation in the rock does not become stable even 220 days after the fully-mechanized face has moved forward. The deformation at a roadway crossing involving three roadways is larger than at the

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