The effect of joint roughness coefficient (JRC) and joint...

International Research Journal of Applied and Basic Sciences © 2013 Available online at www.irjabs.com ISSN 2251-838X / Vol, 4 (8): 2216-2224 Science Explorer Publications

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The effect of joint roughness coefficient (JRC) and joint compressive strength (JCS) on the

displacement of tunnel

Fereydoon Amanloo*1 ,Vahid Hosseinitoudeshki2

1. Department of Civil Engineering� Zanjan Branch� Islamic Azad University� Zanjan� Iran 2. Department of Civil Engineering� Zanjan Branch� Islamic Azad University� Zanjan� Iran

*Corresponding Author email: [email protected]

ABSTRACT: Shear strength of rock mass, joint surfaces and discontinuities and joint dips have important role in the analysis of the stability of underground constructions which designed and implemented in rock. So understanding the factors affecting the shear strength of discontinuity that separates the blocks makes it seem necessary. One way to estimate the shear strength of the joints and joint behaviour in rocks is using empirical theory of Barton. Joint roughness coefficient (JRC) and joint compressive strength (JCS) are the parameters of this equation. This study used data from Tizhtizhgaran tunnel that have been dug in the shale rocks. In this tunnel modeling, seven dips of joints in 0 , 15 , 30 , 45 , 60 , 75 , 90 angles analyzed using Phase2 software and displacement diagrams are plotted on the sections of tunnel under static load. In addition, the effect of increasing joint compressive strength is evaluated on the joint roughness in around of tunnel. The obtain results show that the joint dip, joint compressive strength and joint roughness coefficient have different influences on the displacement of tunnel. With increasing joint roughness coefficient from 0 to 15, displacement in around of tunnel is decreased and from 15 to 20, the displacement is increased. With increasing the dip of joints from 0 to 90, the values of JCS and JRC have greater effect in the displacement in around of tunnel. Keywords: Joint roughness Coefficient (JRC); Joint Compressive Strength (JCS); Tunnel; Joint Dip.

INTRODUCTION

There is a subtle relationship between the construction projects such as tunnel and rock characteristics

and assess the risks of geological phenomena and their effects on the project. These studies are based on library research, field surveys, laboratory and field experiments. The following is an overview of the work done on the effects of structural and physical behaviour of rock joints in the stability of tunnel section. The undulations and asperities on a natural joint surface have a significant influence on its shear behaviour. Generally, this surface roughness increases the shear strength of the surface, and this strength is extremely important for stability of excavation in rock. Patton (1966) demonstrated this influence by means of an experiment in which he carried out shear tests on '' saw – tooth'' specimens such as the one illustrated in Figure1. Shear displacement in these specimens occurs as a result of the surfaces moving up the inclined faces , causing dilation (an increase in volume) of the specimen. The shear strength of patton's saw – tooth specimens can be represented by:(1)

Where bφ is the basic friction angle of the surface and i is the angle of the saw – tooth face.

Figure 1.Patton's experiment on the shear strength of saw – tooth specimens

Intl. Res. J. Appl. Basic. Sci. Vol., 4 (8), 2216-2224, 2013

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Bartton's equation is valid at low normal stresses where shear displacement is due to sliding along the

inclined surfaces. At higher normal stresses, the strength of the intact material will be exceeded and the teeth will tend to break off, resulting in a shear strength behaviour which is more closely related to the intact material strength than to the friction characteristics of the surfaces.

While patton's approach has the merit of being very simple, it does not reflect the reality that changes in shear strength with increasing normal stress are gradual rather than abrupt. Bartton studied the behaviour of natural rock joints and proposed that above equation could be rewritten as: (2)

Where JRC is the joint roughness coefficient and JCS is the joint wall compressive strength. Bartton and choubey (1977) provided the first non–linear strength criterion for rock joints on the basis of

their direct shear test results for 130 samples of variably weathered rock joints.

(3)

Where rφ is the residual friction angle.

Mahendra et al. (2008) also have studied the jointed and highly anisotropic rock masses in the underground construction. It has been shown in this study that the ratio of lateral to axial strain may be very high, especially, if the joints are critically oriented. The assumption of isotropic linearly elastic material is not applicable in such situations. This observation is based on the outcome of an extensive laboratory testing program, in which a large number of specimens of a jointed rock mass with various joint configurations were tested under uniaxial loading conditions. The trends of experimental results for both lateral strain ratio and rock mass strength have also been verified through distinct element modeling. The reason for high lateral strains has been attributed to the creation of voids and also to the fact that permanent deformations due to slip commence along rock joints right from the start to loading process. A simple mechanistic model has been suggested to explain the high values of lateral strain for rough and dilatant rock joints.

Du et al. (2011) have studied the comparison between empirical estimation and direct shear test to measure the joint shear strength in rock. Comparison results show that for natural rock joints with joint surfaces closely matched, the average relative error of joint shear strength between empirical estimation and direct shear test is 9.9 percent. However, for natural rock joints surfaces with joint surface mismatched, the average relative error of joint shear strength between empirical estimation and direct shear test is 39.9 percent.

Prudencio et al. (2007) in laboratory tests on artificial rock models with non – persistent joints illustrated the large anisotropy in the strength of a fractured rock mass. The stress orientation relative to the orientation of the joints and the value of the confining stress resulted in different failure model samples with steeply dipping non – persistent joints and joint step angle larger than 90 underwent planar failure The strength of some samples tuned out to be larger than the strength predicted by a simple model because the normal stress on the rock bridges is several times larger than the stress as summed by the simple model. These tests showed that there are three kinds of failure, planar failure, stepped failure and rotational failure, planar failure and stepped failure are associated with higher strengths, while rotational failure is usually associated with a very low strength, ductile behaviour and large deformation. Rotational failure would lead to regressive slope failure.

Nakagawa et al. (2004) emphasized to the appropriate modeling of mechanical behaviour of discontinuity and its material properties (I – e normal stiffness, shear stiffness, cohesion, friction and dilation angles) so as to evaluate the stability and deformational behavior of structures in the discontinuous rock masses. This paper presented a rational procedure for determination of deformational characters and strength of natural rock joints by using the automated servo – controlled direct shear apparatus. If we apply the technique shown in this paper, we can rationally determine the mechanical properties of natural rock joints. This contributes to improvement of reliability of predicting behaviour of structures in discontinuous rock mass by using discontinuous numerical approach. Han and Tang (2010) have investigated the numerical simulation for anisotropy of compressive strength of rock mass with multiple natural joints. Joints have great influence on compressive strength of jointed rock mass. In this paper based on Rock Failure Process Analysis model RFPA,

[ ])(logtan 10

n

bn

JCSJRC

σφστ +=

[ ])(logtan 10

n

rn

JCSJRC

σφστ +=


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the evolutionary processes of failure process of rock mass with multiple natural Joints were simulated. Panthi and Nilsen (2007) in their research have studied the uncertainty analysis in the physical parameters weak rocks such as shale, phyllite and schist in Himalaya zone before and after excavation. Because the rock mass of weakness / excavation. Fault zones are incapable of sustaining high tangential stress.

MATERIALS AND METHODS

The numerical method using the computational code (phase2) has been applied in analyzing the sections of tunnel. Phase2 is a two – dimensional program which planned based on infinite elasto-plastic element that used for calculation the stresses and displacements around the underground excavations. In this paper tunnel which is used in order to numerical simulation includes Hamro tunnel that excavated in the jointed shale rock in depth of 41 m and diameter of 12 m in the northwest of Iran. Numerical analysis based on two dimensional analyzing and plane strain. The modeling has been with parallel joints in seven dips of joint (0, 15, 30, 45, 60, 75 and 90 degree). In addition, the values of JCS 2 , 5 , 10 , 15 , 20 , 25 and JRC 5 , 10 , 15 , 20 has been analyzed.

Figure 2. Geometric model of the tunnel sections from left to right for

dips of joints in 0 �� , 15 �� , 30 �� , 45 �� , 60 �� , 75 �� , 90 ��

Rock mass in tunnel section The study area is related to the jointed shale rock with the following mechanical properties. The properties of

rock mass including the strength of rock (�cm) deformation, modulus of rock (Em) and constants of rock (mb ,s , a ) have been calculated by Roclab software. This software is provided by Hoek et al. (2002). In this software the strength of rock and deformation modulus are calculated by means of Hoek's equations. In addition, the constants are determined by means of geological strength index (GSI), the intact rock parameter (mi) and the disturbance factor (D) that associated with existing disturbance as a result of excavation. Finally, shear strength rock mass parameters (� , c) are obtained with comparison Mohr–coulomb and Hoek – Brown criterion . The results are shown in Figure 3 and Table 1.

Table 1.Geomechanical properties of rock mass Roclab program's input and output Hoek Brown Classification Hoek Brown Criterion

�ci (Mpa) GSI mi D mb s a

Intact Uniaxial Compressive Strength

Pick GSI Value

Pick MI Value Disturbance Factor D

Hoek-Brown Criterion

25 33 6 0.8 0.111 0.000039 0.518 Mohr-Coulomb Fit Rock Mass Parameters C (Mpa) �(degree) �t (Mpa) �c (Mpa) �cm (Mpa) E dm(Mpa)

Cohesion Friction angle

Tensile strength

Uniaxial compressive strength

Global strength Deformation modulus

0.086 27.38 -0.009 0.130 0.989 1127.30

Intl. Res. J. Appl. Basic. Sci. Vol., 4 (

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Analysis of results

and joint


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Analysis of results In this part of paper the displacement diagrams of tunnel roof in different

and joint compressive strength


Analysis of results In this part of paper the displacement diagrams of tunnel roof in different

compressive strength

Figure 4

Intl. Res. J. Appl. Basic. Sci. Vol., 4 (8), 2216

In this part of paper the displacement diagrams of tunnel roof in different compressive strength (JRC) that resulted from numerical analysis are as follow:

Figure 4. Displacement diagram of tunnel roof in the

2216-2224, 2013

Figure 3. Rock mass parameters

In this part of paper the displacement diagrams of tunnel roof in different that resulted from numerical analysis are as follow:

Displacement diagram of tunnel roof in the

2013

Rock mass parameters

In this part of paper the displacement diagrams of tunnel roof in different that resulted from numerical analysis are as follow:

Displacement diagram of tunnel roof in the

Rock mass parameters

In this part of paper the displacement diagrams of tunnel roof in different joint that resulted from numerical analysis are as follow:

Displacement diagram of tunnel roof in the dip of joint

joint roughness coefficientthat resulted from numerical analysis are as follow:

dip of joint 0º

roughness coefficient

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roughness coefficient (JRC)


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Figure 5. Displacement diagram of tunnel roof in the dip of joint 15

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The above diagrams show that when the values of joints roughness coefficient increases of 0 to 5, the trend of displacement is completely decreasing and displacement of tunnel in all dip of joints especially dips in 30, 45, 60, 75, 15.Decreases and in values from 5 to 15 in the above slopes , the trend of displacement is bowling and the lowest amount of displacement was seen in the roughness coefficient of 10.In the values of 5 to 10 the trend of curve is decreasing with slow movement and in the values of 10 to 15 the trend of curve is Ascending with slow movement .In the roughness coefficient. In the value of 15 to 20 , the trend of curve with a sharp slope is ascending and with increasing the displacement. The stability of tunnel decreases when we investigate the diagrams and considered the influences of joints compressive strength the dependency between two parameters, roughness coefficient and compressive strength can be seen. The above diagrams indicate that in the low roughness coefficient (JRC) that is in values of 0 to 10 when the compressive strength


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becomes higher than 5 to 25 MP0 the trend of carves is in terms of each other , but when the roughness of surfaces and the saw – tooth increase , against the movement trend in the values of 0 to 10 , (JRC) the trend of curves becomes completely ascending , so that of higher compressive strength (JCS) the trend of curve becomes more ascending and the stability of tunnel decreases. Shear displacement of these specimens occur as a result of upward movement of inclined faces, the dilation and swelling occur. One of the reason for the increasing of displacement of rock mass joints is the increasing of roughness coefficient (JRC) and compressive strength values (JCS) . In order to create a necessary condition on the joint rock surface for generating dilation sliding, first we should know that sliding occurs at low normal stresses, and the existing normal stress (� n) should be lower than the stress (� n , crit) (shear normal stress of joint surface) otherwise , sow–tooth surfaces of joints crack and the dilation doesn't occur (Figure 11).Second if existing shear stress is higher than existing stress on the planar surfaces the compressive strength continuously increases and the rock dilates. So because in higher roughness coefficients, the saw tooth surfaces become larger and rougher and with increasing the compressive strength, the saw–tooth stay intact and their strength against cracking decreases , the only alternative for jointed rock mass is that at low normal stress acts as dilation (such as figure 12) and increases the displacement meanwhile in higher roughness coefficient, the saw – tooth surfaces become larger and rougher, so dilation in these coefficients is larger and also in some surveys we observed that with increasing the roughness coefficient inspite of what has said. The displacement decreases that this is as a result of joints low compressive strength. This causes the lower face of joint cracked and this occurs in low compressive strengths of 5 to 15 Mpa.

Figure 11.Profile for finding the joints with low compressive strength

Figure 12.Dilation and sliding of joints on each other and distancing the upper surface.

Also based on Barton's empirical model we have: (4)

In the Barton's model the underlined expression indicates the dilation angle that is shown with the symbole of dilation angle 5. It shows the dilation angle ratio in joint shear strength. Due to the above formula we concluded that as the values of roughness coefficient increase, the dilation angle increases too. It indicates that the influence that roughness coefficient has on the joint in higher coefficient is more than the influence of joint compressive strength. and in higher roughness. dilation is larger, Also we observed that the existing normal stress is in the root of the fraction , so as its value becomes less , the stress on the upper layers on the

[ nnp J R Cσστ += )(l o g.t a n1 0


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joint surface inclines and the movements of joints on each other increase . Based on equation 4, as the dip of joints increase of 0

º to 90

º, existing normal strews inclines and the surfaces of joints instead of cracking on each

other, distant along the inclined faces of joints and displacement increase. (5) (Normal stress in situ on a surface inclined by �)

CONCLUSION The obtain results from numerical analysis of excavation of tunnel in the jointed rock mass as follows: With increasing joint roughness coefficient (JRC) from 0 to 15, displacement in around of tunnel is decreased and stability of tunnel is increased. With increasing joint roughness coefficient (JRC) from 15 to 20, displacement in around of tunnel is increased. Dilation has increased displacement in around of tunnel as the roughness values increase. In lower roughness coefficient, the compressive strength has greater role in the stability of joints and displacement in around of tunnel is decreased. In higher roughness coefficient, the compressive strength has greater role in the dilation of joint surfaces and displacement in around of the tunnel is increased. In higher dips of joints with decreasing of the normal stress, the dilation of joints increases. With increasing the dip of joints from 0

º to 90

º, the values of JCS and JRC have greater effect in the

displacement in around of tunnel.

REFERENCES Barton NR, Choubey V. 1977. The shear strength of rock joints in theory and practice, Rock Mech, 10(1-2): 1-54. cases from Nepal Himalaya, International Journal of Rock Mechanics & Mining Sciences, 44(2007) : 67–76. Du S, Hu Y, Hu X, Guo X. 2011.Comparison between Empirical Estimation by JRC-JCS Model and Direct Shear Test for Joint Shear

Strength, Journal of Earth Science, 22(3): 411–420. Han F, Tang C. 2010. Numerical investigation for anisotropy of compressive strength of rock mass with multiple natural joints, Journal of

coal science and engineering,16: 246–248. Nakagawa M, Jiang Y, Kawakita M, Yamada Y, Akiyama Y.2004. Evaluation of mechanical properties of natural rock joints for

discontinuous numerical analysis.Proceeding of the ISRM International Symposium 3rd ARMS. Ohnishi and Aoki (eds). Panthi KK, Nilsen B.2007. Uncertainty analysis of tunnel squeezing for two tunnel Patton FD. 1966. Multiple modes of shear failure in rock. Proc. 1st congr. Int. Soc. Prudencio M, Van Sint Jan M.2007. Strength and failure modes of rock mass models with non-persistent joints, International Journal of

Rock Mechanics & Mining Sciences, 44(2007) ,890–902. Rock Mech., Lisbon 1: 509-513. Singh M, Singh B.2008. High lateral strain ratio in jointed rock masses,Engineering Geology, 98(2008): 75–85.

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0 coshn =

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