Engineering geological appraisal of the rock masses and preliminary support design, Dorukhan Tunnel, Zonguldak, Turkey

Engineering Geology 92 (2007) 14 – 26 www.elsevier.com/locate/enggeo

Engineering geological appraisal of the rock masses and preliminary support design, Dorukhan Tunnel, Zonguldak, Turkey M. Genis a , H. Basarir b,⁎, A. Ozarslan a , E. Bilir a , E. Balaban c a

Zonguldak Karaelmas University, Engineering Faculty, Mining Eng. Department, 67100 Zonguldak, Turkey b Inonu University, Engineering Faculty, Mining Eng. Department, 44280 Malatya, Turkey c Turkish General Directorate of Highways, 15th District, 37100 Kastamonu, Turkey Received 14 November 2006; received in revised form 22 February 2007; accepted 28 February 2007 Available online 12 March 2007

Abstract This paper presents the results of engineering geological studies of the rock masses along a road tunnel. Rock mass qualities of the rock units along the tunnel were determined by means of Rock Mass Rating (RMR), Geomechanic Classification (Q) system, Geological Strength Index (GSI), Rock Mass Index (RMi) and New Australian Tunneling Method (NATM). In order to determine tunnel stability, necessary support types and categories RMR, Q, RMi and NATM systems were employed as empirical tunnel support design methods. However, these empirical design guidelines for tunnel support based on rock mass classification systems failed to analyze the support performance. The performances of the proposed support systems were analyzed by means of numerical analysis, described in this paper. A 2D finite element analysis program was used as numerical method. The necessary rock mass parameters were obtained by means of rock mass classification systems. © 2007 Elsevier B.V. All rights reserved. Keywords: Rock mass classification system; Engineering geology; RMR; Q; GSI; RMi; NATM; Numerical method; Tunnel support design

1. Introduction The Yenicaga–Zonguldak highway is one of the most important highways connecting the Black sea coast to the Central Anatolian region. Due to the geotechnical problems associated with the existing tunnel and heavy traffic load, construction of a new tunnel called the Dorukhan Tunnel, was decided by General Directorate of Highways authorities. The location of project area is given in Fig. 1. The planned length of the tunnel is 1030 m and it will accommodate two traffic lanes. The driven tunnel

⁎ Corresponding author. Tel.: +90 422 3410010/4507; fax: +422 3410046. E-mail address: [email protected] (H. Basarir). 0013-7952/$ - see front matter © 2007 Elsevier B.V. All rights reserved. doi:10.1016/j.enggeo.2007.02.005

will have a modified ellipsoid shape with excavated dimensions of 11.8 m width and 8.3 m height (Fig. 2). Engineering geological studies and rock mechanics experiments were conducted both in the field and the laboratory. The field studies included geological mapping, core drilling, discontinuity surveying and geotechnical descriptions. Rock mass classification systems are very useful tools for the preliminary design stage of a project, when very little detailed information on rock mass are available. On the other hand, utilization of several rock mass classification systems is recommended to build up a picture of composition and characteristics of rock mass to provide initial estimates of support requirements (Hoek et al., 1995). The tunnel stability and the required support systems were assessed by

M. Genis et al. / Engineering Geology 92 (2007) 14–26

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Fig. 1. Location map of the Dorukhan Tunnel.

means of RMR, Q, GSI, RMi and NATM rock mass classification systems. Although rock mass classification systems are very useful during the preliminary design stage, they cannot adequately calculate stress distributions, support performance and deformations around the tunnel. Therefore empirical methods should be augmented by numerical methods (Kontogianni and Stiros, 2002; Sari and Pasamehmetoglu, 2004; Basarir, 2006). Since rock mass strength parameters are essential input parameters for the numerical methods, a number of studies were performed to estimate these parameters by means of rock mass classification systems. In this study, rock mass strength parameters were obtained by means of RMR, Q, GSI and RMi systems.

Ordovician: Bolu Granitoid (Erendil et al., 1991) Quaternary aged Alluvium composed of sandy gravel. The Devonian aged Aksudere formation is represented by mainly phyllite, shale, recristalized limestone, dolomitic limestone (Erendil et al., 1991). Phyllite shows well-developed schistosity and the main mineralogical constituents are quartz, sericite, muscovite, chlorite, epidote and feldspar. They also host some embedded massive recristalized limestone blocks. Bolu Granitoid is composed mainly of granodiorite and granite (Erendil et al., 1991). Bolu Granitoid is greenish gray in color, coarse-grained and their mafic content ranges from 10% to 15%. Diabase and quartz dykes, generally located along faults, are associated with the Bolu Granitoid.

2. Geology 3. Engineering geology The study area is located in the Western Pontides. The geological formations mainly consist of igneous and sedimentary rocks. The geology consists of the following formations: Quaternary: Alluvium deposit Devonian: Aksudere Formation (Erendil et al., 1991)

The engineering geological studies include both field and laboratory studies. The field studies consist of field observation, boreholes and discontinuity surveys. Laboratory tests were conducted on samples, collected from the field and the boreholes. A geological crosssection along the tunnel is given in Fig. 3.

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Fig. 2. The Dorukhan Tunnel cross-section.

The main rock types along the tunnel alignment include phyllite, tectonic breccia and moderately weathered granodiorite. Phyllite is a member of the Aksudere formation. Following the intrusion of Bolu granodiorite, breccia result from tectonic faulting occurred. Tectonic breccia composed of phyllitic and granitic fragments in the clayey–silty mylonitic matrix. Phyllite is bluish-greenish gray in color, lens shaped, banded and moderately weathered. The uniaxial compressive strength (UCS) class of phyllite is medium with average strength of 30 MPa. The average RQD of phyllite is 26%. Joints are moderately spaced and range from 20 cm to 30 cm. Joints are highly persistent and their surfaces are planar and slightly rough. Apertures are 5 mm in width filled by silty sand. A very favorable discontinuity orientation is observed. The heavily broken phyllite is the most problematic rock unit along the tunnel alignment. This unit is in bluishgreenish gray color and it is moderately to highly weathered. This rock unit has very weak strength with

the average UCS being 1.5 MPa. The average RQD for this rock unit is 10%. Joints are very closely spaced and in the range from 20 to 60 mm. They show high persistency and their surfaces are undulating smooth. Apertures are mostly bigger than 5 mm and are filled by silty clay. The tectonic breccia has weak strength with an average UCS of 15 MPa. It is highly weathered and the average RQD is 59%. Spacing of discontinuities ranges from 6 to 20 mm, which is classified as close spacing. Discontinuity surfaces are slickenslided with clay infilling and possess high persistence. Apertures range from 1 to 5 mm. The moderately weathered granodiorite is gray in color. Average RQD and UCS of this rock unit are 50% and 31 MPa, respectively. Joint spacing ranges from 20 to 30 cm and is classified as close spacing. Highly persistent joints are observed and 2 mm wide apertures are filled with silty sand. In total 149 discontinuities were measured in the field. Discontinuity orientations were processed by computer software DIPS 5.1 (Rocscience, 2002), based on equal-area stereographic projection and dominant

Fig. 3. The Dorukhan Tunnel geological section.

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Fig. 4. Dominant discontinuity sets.

discontinuity sets were distinguished. The determined dominant discontinuity sets are illustrated in Fig. 4. The dip and dip directions of main discontinuities were determined as 371/248, 82/149, 87/002 and 87/355. Laboratory experiments were conducted on core specimens of NX size, 54 mm, taken from core drillings. Laboratory experiments were carried out in accordance with the methods suggested by ISRM (ISRM, 1981) to determine the physical and mechanical properties of rock units, including unit weight, uniaxial compressive strength. Triaxial compressive strength tests were also conducted on core specimens to determine m and s Hoek–Brown constants of intact rock. Table 1 presents the results of the laboratory tests that were performed by the Turkish General Directorate of Highways, Technical Research Department, Soil Mechanics and Tunnel Section.

domain, Jn is the rating for the number of joint sets in the same domain, Jr is the rating for the roughness of the least favorable of these joint sets or filled discontinuities, Ja is the rating for the degree of alteration or clay filling of the least favorable joint set or filled discontinuity, Jw is the rating for the water inflow and pressure effects, which may cause outwash of discontinuity infillings, and stress reduction factor (SRF) is the rating for faulting, for strength/stress ratios in hard massive rocks, for squeezing or for swelling (Barton, 2002). A stress free form of Q was defined later by Goel et al. (1995) as QN. In order to calculate QN, stress reduction factor (SRF) is taken as 1, which is given in Eq. (2):

 RQD Jr QN ¼ ð2Þ Jw Jn Ja

4. Rock mass classification systems

Hoek et al. (1995) proposed the modified Tunneling Quality Index, Q′, calculated in the same way as the

In this paper, RMR (Bieniawski, 1989), Q (Barton et al., 1974; Grimstad and Barton, 1993), GSI (Hoek et al., 1995), RMi (Palmström, 2000) rock mass classification systems were employed to characterize the rock mass and to estimate the rock mass strength parameters. The Q-value is estimated from the following expression (Barton et al., 1974): Q¼




 RQD Jr Jw Jn Ja SRF

ð1Þ

where RQD is the percentage of competent drill-core sticks N100 mm in length (Deere et al., 1967) in a selected

Table 1 Physical and mechanical properties of the rock materials Parameters, symbol, unit

Phyllite Heavily broken Tectonic Moderately phyllite breccia weathered granodiorite

Uniaxial compressive strength, UCS, MPa Unit weight, γ, t/m3 mi constant si constant

30

1.5

15

30.8

2.85

1.96

2.08

2.68

7 1

7 1

19 1

26 1

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Table 2 The estimated rock mass classification systems Formation

RMR

Q

GSI

RMi

Qc

QN

Q′

Phyllite Heavily broken phyllite Tectonic breccia Moderately weathered granodiorite

29–34 (32) 12–19 (16) 25–29 (27) 39–43 (41)

0.07–0.1 (0.09) 0.01–0.02 (0.02) 0.06–0.09 (0.08) 1.04–1.38 (1.21)

25–30 (28) 10–15 (13) 25–30 (28) 35–40 (38)

0.105–0.145 (0.125) 0.002–0.004 (0.003) 0.047–0.075 (0.061) 0.351–0.463 (0.407)

0.02 0.00 0.01 0.37

0.60 0.15 0.20 1.21

0.91 0.30 0.61 1.83

Table 3 The proposed empirical equations for calculation of Emass Researcher

Equation no.

Equation

Bieniawski (1978)

(5)

Emass = 2RMR − 100 (GPa)

Notes ðRMR−10Þ 40

Serafim and Pereira (1983)

(6)

Emass ¼ 10

Grimstad and Barton (1993)

(7)

Emass = 25 logQ (GPa)

(GPa)

0.375

Palmström (2000)

(8)

Hoek and Brown (1998)

(9)

Read et al. (1999)

(10)

RMi (GPa) Emass = 5.6 rffiffiffiffiffiffiffi ffi rci ðGSI−10 10 40 Þ (GPa) 100  3 (GPa) Emass ¼ 0:1 RMR 10

Barton (2002)

(11)

Emass ¼ 10Qc (GPa)

Emass ¼

For RMR N 50 For RMR b 50 For Q N 1 For RMi N 0.1 For σci b 100 MPa

1=3

parameter Qc has been defined by Barton (2002) as below:

standard Q rock mass classification, except that the stress reduction factor (SRF) and joint water reductions factor (Jw) was set to 1.00.

 RQD Jr Q V¼ ð3Þ Jn Ja

Qc ¼ Q

rci 100

ð4Þ

where σci is the strength of intact rock in MPa. RMR, Q, GSI, RMi, QN, Qc and Q′ values are presented in Table 2. In order to overcome some of the

In 2002, the Q system was re-compiled to improve correlation between engineering parameters and a new

Table 4 The proposed empirical equations for calculation of σcmass Researcher

Equation no.

Equation

Hoek and Brown (1980) Yudhbir et al. (1983) Ramamurthy (1986) Goel (1994)

(12)

rcmass

(13) (14)

pffiffiffiffiffiffiffiffiffiffiffiffiffiffiffi ðRMR−100Þ ¼ rci e 9 (MPa) ðRMR−100Þ 7:65 100

rcmass ¼ rci e r

ðRMR−100Þ cmass¼rci e 18:75

(MPa)

5:5gQV1=3 (MPa) B0:1

rcmass ¼

Kalamaris and Bieniawski (1995) Palmström (2000) Bhasin and Grimstad (1996) Sheorey (1997) Trueman (1998)

(16)

r

(17) (18)

σcmass = RMi P (MPa)  rci = σciJ1=3 rcmass ¼ 100 7gQ (MPa)

(19) (20)

Aydan and Dalgic (1998) Barton (2000) Hoek et al. (2002)

(21)

rcmass ¼ rci e 20 (MPa) rcmass ¼ 0:5e0:06RMR (MPa) RMR rci (MPa) rcmass ¼ RMR þ bð100−RMRÞ  rci 1=3 rcmass ¼ 5g Q 100 (MPa) rcmass ¼ rci sa (MPa)

(22) (23)

ðRMR−100Þ 24

σci is the strength of intact rock (MPa)

(MPa)

(15)

cmass¼rci e

Notes

γ is the density of rock mass (t/m3)

(MPa)

γ is the density of rock mass (t/m3)

RMR−100

β=6

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Table 5 The proposed empirical equations for calculation of m and s contants of rock mass Researcher

Equation no.

Equation

Hoek et al. (1995) Hoek et al. (1995) Palmström (2000) Palmström (2000) Hoek et al. (2002) Hoek et al. (2002)

(24) (25) (26) (27) (28) (29)

¼ 0:135ðQVÞ1=3 s = 0.002Q′ s = JP2 m = mi JP0.64 GSI−100 m eð 28−14D Þ mi ¼ GSI−100 s¼eð 9−3D Þ m mi

uncertainties of the classification systems, a range of rock mass values was estimated rather than just a single value. 5. Estimating rock mass properties Rock mass properties such as Hoek–Brown constants, deformation modulus of the rock masses and uniaxial compressive strength of the rock mass were calculated by using empirical equations based on Qc, QN, Q, RMR, RMi and GSI. 5.1. Deformation modulus of rock masses In-situ determination of the deformation modulus of rock mass (Emass) is costly and often very difficult. Thus, empirical methods are generally used in estimating Emass (Basarir et al., 2005). By means of the empirical methods, Emass can be easily acquired. The proposed equations by different researchers are presented in Table 3. 5.2. Strength of rock masses Different researchers have proposed different empirical equations to calculate the strength of rock mass (σcmass) based on rock mass classification systems. The most widely used equations are tabulated in Table 4. Since the strength value, obtained from Eq. (15), is too high when compared to the other strength values, this value was not used in calculating the average strength value of the rock masses.

techniques were applied and thus the value of D was considered to be zero. The calculated deformation modulus, strength and Hoek–Brown constants of rock masses for the present work are given in Table 6. The averages of these parameters are calculated using the weighted average method and used in the numerical modeling. 6. Empirical support design Based on Q, RMR, RMi and NATM classification systems, the necessary support systems and proposed excavation methods for the rock units along the tunnel route are presented in Table 7. The New Austrian Tunneling Method (NATM) classes of rock units were determined by using correlations with the RMR and Q system according to the procedure given by the Turkish General Directorate of Highways (1997). Considering the support and excavation method recommendations of empirical methods, two different

Table 6 Calculated rock mass strength parameters Parameter

Emass, GPa

Phyllite

Heavily broken phyllite

Moderately weathered granodiorite

Breccia

(5) (6) (7) (8) (9) (10) (11)

– 3.55 – 2.57 1.54 3.28 2.88 2.76 0.69 0.17 0.80 11.86 1.76 0.13 2.58 1.00 3.41 2.18 4.11 0.45 1.57 0.92 0.21 0.54 0.55 1.8 E -3 1.7 E -5 0.3 E -3 0.7 E -3

– 1.41 – 0.63 0.15 0.41 0.67 0.65 0.01 0.00 0.02 5.64 0.05 0.00 0.06 0.02 1.31 0.05 0.66 0.00 0.20 0.63 0.13 0.31 0.36 0.6 E -3 0.4 E -5 – 0.3 E -3

– 5.96 2.07 4.00 2.78 6.89 7.20 4.82 1.16 0.34 1.32 14.09 2.64 0.41 6.16 1.61 5.85 3.19 9.64 0.90 3.02 4.30 1.63 2.84 2.92 3.7 E- 3 1.8 E- 4 1.0 E- 3 1.6 E- 3

– 2.66 – 1.96 1.09 1.97 2.29 1.99 0.26 0.06 0.31 7.56 0.72 0.06 0.94 0.39 2.53 0.87 2.38 0.22 0.79 2.17 0.56 1.45 1.39 1.2 E -3 1.7 E -5 0.3 E -3 0.5 E -3

Average

σcmass, MPa

5.3. Hoek–Brown parameters Hoek et al. (1995), Palmström (2000) and Hoek et al. (2002) suggested some empirical equations to calculate m and s parameters of rock masses mm, sm. The suggested equations are given in Table 5. In Hoek et al. (2002) equation D is the disturbance factor that depends on the amount of disturbance in the rock mass associated with the method of excavation (e.g. smoothness of blasting). In this study it was assumed that blasting quality was excellent and controlled blasting

Equation no.

(12) (13) (14) (15) (16) (17) (18) (19) (20) (21) (22) (23)

Average m constant

(24) (27) (28)

Average s constant Average

(25) (26) (29)

M. Genis et al. / Engineering Geology 92 (2007) 14–26

excavation methods and support types were proposed for the rock units along the tunnel route. Tunnel sections, the rock units, proposed excavation methods, round length, support systems and support installation time are given in Table 8. The proposed excavation methods and support types are also illustrated in Figs. 5 and 6. 7. Numerical modeling The objective of numerical modeling is to check the validity of the proposed support systems and excavation methods given in Table 8. The computer software Phase2, a 2D Finite Element Program developed by Rocscience (1998), was used for

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calculating the stresses, the deformations and the thickness of the developed plastic zone around tunnel. The software permits two-dimensional study of the nonlinear deformations of rocks using Hoek–Brown failure criterion. The input parameters presented in Table 6 are uniaxial compressive strength, Young's modulus, Poisson's ratio and m and s Hoek–Brown constants of rock masses. In this program, an automatic mesh around the tunnel is generated and based on the elasto-plastic analysis, deformations and stresses are computed. For modeling tunnels driven in rock units along the tunnel alignment, different finite element models were generated. In all models tunnel geometry and horizontal to vertical stress ratio are the same. Tunnel width and height

Table 7 Empirical tunnel support categories and excavation methods Rock unit

Phyllite

Heavily broken phyllite

Tectonic breccia

Moderately weathered granodiorite

Q Span/ESR 0.07–0.1 (0.09) support 11.8 150 mm thick steel fiber reinforced shotcrete (SFRS). 4 m long 1.7 m spaced rock bolts. RMR 29–34 (32) support Systematic bolts 4–5 m long, spaced 1–1.5 m in crown and walls with wire mesh. Shotcrete: 100–150 mm in crown and 100 mm in sides. Light to medium ribs spaced 1.5 m where required. Excavation Top heading and bench 1.0–1.5 m advance in top heading. Install support concurrently with excavation, 10 m from face.

0.01–0.02 (0.02) 11.8 250 mm thick steel fiber reinforced shotcrete (SFRS). 4 m long 1.5 m spaced rock bolts.

0.06–0.09 (0.08) 11.8 150 mm thick steel fiber reinforced shotcrete (SFRS). 4 m long 1.7 m spaced rock bolts. 25–29 (27) Systematic bolts 4–5 m long, spaced 1–1.5 m in crown and walls with wire mesh.

Shotcrete: 150–200 mm in crown, 150 mm in sides, and 50 mm on face. Medium to heavy ribs spaced 0.75 m with steel lagging and forepoling if required. Close invert. Multiple drifts 0.5–1.5 m advance in top heading. Install support concurrently with excavation. Shotcrete as soon as possible after blasting.

Shotcrete: 100–150 mm in crown and 100 mm in sides. Light to medium ribs spaced 1.5 m where required. Top heading and bench 1.0–1.5 m advance in top heading. Install support concurrently with excavation, 10 m from face.

1.04–1.38 (1.21) 11.8 90 mm thick steel fiber reinforced shotcrete (SFRS). 4 m long 2.5 m spaced rock bolts. 39–43 (41) Systematic bolts 4 m long, spaced 1.5–2 m in crown and walls with wire mesh in crown. Shotcrete: 50–100 mm in crown and 30 mm in sides.

RMi support 0.105–0.145 (0.125) 5 m length rock bolts spaced 1 m. 150–250 mm thick fiber reinforced shotcrete. NATM B3 Class Utilization of systematic support support and local forepoles.

0.002–0.004 (0.003) Short blast round, shotcrete quickly after blast and concrete lining.

0.047–0.075 (0.061) Special designed shotcrete or concrete lining.

C2 Utilization of systematic support.

B3 Utilization of systematic support and local forepoles.

Excavation

Use Road header. Multiple drifts, 1.0–1.5 m advance in top heading and 2 m advance in bench.

Use smooth blasting or road headers where necessary. Top heading (1.5–2.0 m) and bench (2.5 m).

Use smooth blasting or road headers where necessary. Top heading (1.5–2.0 m) and bench (2.5 m).

12–19 (16) Systematic bolts 5–6 m long, spaced 1–1.5 m in crown and walls with wire mesh. Bolt invert.

Top heading and bench 1.5–3 m advance in top heading. Commence support after each blast. Complete support 10 m from face. 0.351–0.463 (0.407) 4.5 m length rock bolts spaced 1.25 m. 100–150 mm thick fiber reinforced shotcrete. B2 Utilization of systematic support and local forepoles if necessary. Use smooth blasting. Top heading (2.0–2.5 m) and bench (3.5 m).

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Table 8 Proposed excavation methods and support types for the rock units along the tunnel route Tunnel section

Rock unit

Excavation method

Round length

Support system

29 + 700/29 + 826 29 + 962/29 + 990 29 + 990/30 + 058 30 + 058/30 + 154 29 + 929/29 + 962 30 + 154/30 + 730

Phyllite

Top heading and bench 1.0–1.5 m advance in top heading.

1–1.5 m

Install support concurrently Systematic bolts 4–5 m long, spaced 1–1.5 m in crown and walls with wire mesh. Shotcrete with excavation, 1.50 m 100–150 mm in crown and 100 mm in sides. from face. Light to medium ribs spaced 1.5 m where required.

29 + 826/29 + 863 29 + 863/29 + 929 30 + 058/30 + 109

Breccia Moderately weathered granodiorite Heavily Multiple drifts 0.5–1 m broken 0.5–1.5 m phyllite advance in top heading.

Systematic bolts 5–6 m long, spaced 1–1.5 m in crown and walls with wire mesh. Bolt invert. 150–200 mm in crown, 150 mm in sides, and 50 mm on face. Medium to heavy ribs spaced 0.75 m with steel lagging and forepoling if required. Close invert.

are 11.8 and 8.4 m, respectively. The tunnel lies at a relatively shallow depth, the depth of the tunnel in generated models changes from 60 to 150 m depending on the depth of the modeled rock unit. The outer model boundary was set to be at a distance of 10 times of the radius of tunnel. 3555 three-noded-triangular elements were used in the mesh. Finer zoning was used around the excavation. The Hoek–Brown failure criterion was used to estimate the yielded elements and the plastic zone of rock masses in the vicinity of tunnel. Rock mass behavior was represented as an elastic perfectly plastic material in which failureinvolving slip along intersecting discontinuities as is assumed to occur with zero plastic volume change

Distance between face and support

Install support concurrently with excavation. Shotcrete as soon as possible after blasting.

(Duncan-Fama, 1993) in finite element analysis. This approach is considered as valid for the following reasons: In this study the rock masses are jointed such that spacing are small compared to the size of opening, there are sufficient numbers of joint sets to assure isotropic strength properties for the rock masses and there is not any particular joint set dominating the behavior of the rock masses. The rock mass properties assumed in finite element analysis are obtained from the estimated values given in Section 5. Vertical stress is a function of overburden. It is more difficult to estimate horizontal stress, σh. It is known that they are variable at shallow depth, tending to a hydrostatic state in deep environment (Hoek and Brown, 1978). In

Fig. 5. Proposed excavation sequence and support elements for Phyllite, Breccia and Moderately weathered granodiorite.

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Fig. 6. Proposed excavation sequence and support elements for heavily broken Phyllite.

this research the ratio of horizontal to vertical stresses (k) is assumed to be 1 as suggested by Hoek (2003). Both unsupported and supported cases were analysed for each rock unit. First of all unsupported cases were modelled. Then support application to the excavation boundary was modelled for each case. The support elements used in all models are composed of rock bolts and shotcrete as proposed by the empirical methods. The properties of support elements, such as length, pattern of bolts and thickness of shotcrete are same as those proposed in Table 8. Phase2 model applies support immediately after the excavation. However in real cases some deformation is allowed to occur and installation time of support system takes time, in this time rock mass around tunnel has already shown a certain reduction of stress state. To simulate delayed support installation a load splitting option of the software was used for supported cases. That is, some deformation is allowed to take place between excavation and support installation stages. The enlarged view of excavation sequence and the installation of the support systems for the tunnel driven in heavily broken phyllite are given in Fig. 7. The model for examining the supported case with Phase2 included different; excavation of top heading was achieved and support elements composed of rock bolts and shotcrete were installed, bench excavation was performed, necessary support elements were placed and entire excavation of the tunnel was completed by invert excavation Finally only shotcrete was used to support the invert of tunnel. For unsupported cases, the thickness of the plastic zones, the yielded elements and the maximum total

displacements of the tunnels excavated in different rock types are shown in Fig. 8. As it can be seen from Fig. 8, for the tunnels driven in phyllite, breccia and moderately weathered granite displacements are small. However, the extent of plastic zone and yielded elements suggest that there would be a stability problem for the tunnel. When Fig. 8 is examined, it is more important to consider the extent of plastic zone and yielded elements rather than the magnitude of the displacements. The most problematic formation along the tunnel line is the heavily broken phyllite. Maximum total displacement, lots of yielded elements are observed and larger plastic zone developed around the tunnel as shown in Fig. 8. Therefore, the heaviest support elements are necessary for this formation as presented in Table 8. After support installation, not only the number of yielded elements but also the extent of plastic zone decreased as shown in Fig. 8. For phyllite, moderately weathered granodiorite and tectonic breccia the extent of failure zone decreased significantly, yielded elements are almost disappeared and the total displacement is nearly reduced by two-folds with respect to the induced displacement without support. As for the most problematic heavily broken phyllite formation, the total displacement is 333 mm before support installation is reduced to 18 mm by installing the proposed support system. However, in weaker rock formations such as heavily broken phyllite, some tunneling problems in terms of surprises are expected. These rock masses surrounding the tunnel may be improved by grout injection, and to create protective

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Fig. 7. Numerical modelling stages for Model II.

umbrella ahead of face grouted pipe forepoles can be placed. Also the thickness of the plastic zone around tunnel reduced significantly. This indicates that the proposed support systems were adequate to obtain tunnel stability.

8. Conclusions Based on the information collected from the field and boreholes, the rock mass and material properties were estimated.

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Fig. 8. Displacements and yielded elements for unsupported and supported tunnelling cases.

Rock mass classification systems indicate that some stability problems exists for the rock mass along tunnel route and support measures are necessary. By considering the support recommendations of the empirical methods, support systems and excavation methods were proposed for the rock masses. For the unsupported tunneling cases, stability problems are also verified by the numerical method. Numerical modelling was utilized to evaluate the performance of recommended support system. The ne-

cessary rock properties for numerical modelling were obtained from empirical equations using rock mass classification systems. When the recommended support systems were applied, not only the number of yielded elements but also maximum total displacements were reduced significantly in numerical analysis. The numerical analyses were all performed prior to the start of the tunnel construction and, as such, based on an assessment of rock properties from the engineering geological studies and empirical relationships. Here it

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