Index properties and observations for design of chambers in rock

Engineering Geology, 12 (1978)113--142

113

© Elsevier Scientific Publishing Company, Amsterdam -- Printed in The Netherlands

INDEX PROPERTIES AND OBSERVATIONS CHAMBERS IN ROCK

FOR DESIGN OF

E.J. CORDING and J.W. MAHAR

Department of Civil Engineering, University of Illinois at Urbana-Champaign, Urbana, Ill. 61801 (U.S.A.) (Received April 22, 1977; revised version accepted February 27, 1978)

ABSTRACT Cording, E.J. and Mahar, J.W., 1978. Index properties for design of chambers in rock. In: W.R. Judd (Editor), Near Surface Underground Opening Design. Eng. Geol., 12: 113--142. The in-situ methods emphasized in this paper are used to relate rock index properties to the performance of a tunnel or large chamber. The methods are carried out by exploration prior to construction and by observations during excavation. The mechanics of the deformation and the possible modes of tunnel failure should provide the basis for relating the significant rock index properties to behavior. Index properties of significance in tunneling can be divided into the following categories: the average quality of the rock mass, the properties of planar discontinuities, and the properties of intact specimens. The average quality of the rock mass can be determined from quantitative estimates of rock quality, fracturing, and weathering in the rock core. Calibrated descriptive estimates may suffice. The average rock quality indices are particularly useful in the exploration stage in locating the low quality zones that may cause tunneling difficulties. Properties of planar discontinuities, have a large influence on the performance of large chambers but are more difficult to assess in the exploratory phase of a project. Loosening ground and highly stressed ground conditions and their influence on performance of large chambers are summarized. The following design aspects are discussed: (1) heading support requirements; ( 2 ) support spacing; (3) support displacements; (4) support pressures and bolt lengths; (5) compatibility of initial and final support and integration of excavation and support sequences in large chambers. The size of the chamber with respect to the wave length, continuity, and spacing of joints has an important influence on support pressures and support selection. Classification systems that directly relate rock indices to support can be used for a preliminary estimate of support requirements. However, an additional step should be included when evaluating performance: (1) Evaluate geology and determine significant index properties. (2) Outline expected ground behavior and its effect on construction. This additional step requires a knowledge of construction procedures as well as the models of rock behavior that are appropriate to the given geologic setting and excavation geometry. The geologic data should be plotted in plan and cross-section views of the proposed opening for each excavation sequence in order to identify potential problems and to check the adequacy of the proposed support systems. (3) Select support systems and construction procedures that permit the headings to he safely and economically excavated, and that provide a permanent support suitable over the intended life of the facility.

114 During construction, much can be gained by observing overbreak, fallout, evidence of slabbing due to stress, rock movement, and support distress. The information should be collected at the heading of the tunnel, as the rock is excavated and supported. Observations of support procedures immediately after blasting are particularly important. Rugged, reliable instruments that can be installed, read, interpreted, and related to geologic conditions in a timely manner can contribute much to the observation program.

1. INTRODUCTION The design o f a tunnel or chamber in rock involves n o t only the proportioning o f the final lining but also siting o f the tunnel and selection of tunnel geometry, and selection of excavation and support sequences so t h a t the tunnel can be safely advanced and so that excessive damage to nearby structures is minimized. The geotechnical aspects of the design of a tunnel or chamber are determined from exploration during the initial phases of the project and from observations during and following excavation. During exploration, rock index properties should be determined so that geologic features of significance in tunneling can be evaluated. During construction, observations provide a means of relating rock index properties and construction procedures to the performance of the tunnel. As outlined by Peck et al. (1953), index properties for rock or soil should be quantitative and should be related to the significant engineering properties o f the material. The index p r o p e r t y should be reproducible and should be obtainable from a simple test, preferably from normal e x p l o r a t o r y information. It should be recognized that, at present, many of the index properties in a rock mass are difficult to assess in the e x p l o r a t o r y stage and, even at later stages of the project, are n o t always easily identified or quantified. Moreover, it is even more difficult to predict the ground behavior based on the index properties particularly during the e x p l o r a t o r y phase of the project. The mechanics of the d e f o r m a t i o n and t he possible modes of tunnel failure should provide the basis for relating the significant rock index properties to behavior. Analytical studies, t e m per e d by a knowledge of construction conditions in the field, can be of assistance in developing relations between index properties and tunnel performance. Rock movement, overbreak, and support loads and distress are aspects of the per f or m ance that should be observed. There is a need for developing f ur t he r understanding o f the influence of geologic and construction conditions on the performance of tunnels and chambers. Identification of the i m p o r t a n t index properties as t hey affect rock behavior should be built up f r om extensive observations of geologic conditions, ground behavior, and construction procedures within a given geologic environment. Section 2 o f this paper is an overview o f exploration and index properties for a tunnel or chamber in rock; Section 3 discusses some of the factors influencing stability and support; and Section 4 outlines the observations that can be used to evaluate tunnel performance. Throughout , comparisons are

115

made between methods used for small tunnels, large deep chambers, and shallow chambers in rock. Most of the discussions refer to tunnels and chambers excavated in loosening ground (rock fragments, blocks, and wedges tending to separate from the surrounding rock mass and move under gravity into the opening). Although other ground conditions such as squeezing, swelling, slaking, flowing, and excessive water flows are encountered in tunnels, treatment of these ground conditions is beyond the scope of this paper. 2. INDEX PROPERTIES

2.1. Summary of index properties Rock properties of significance in tunnel design and construction can be categorized in three broad areas: {1) The average quality of the rock mass, a property of the rock mass as distinguished from the properties of an intact rock sample. The quality is a summation of the effect of planar features such as joints and shears on the average properties of the rock mass. The compressibility of the rock mass and, to some extent, the stability of an opening and the tendency of the arch of an opening to ravel are affected by the rock-mass quality. (2} Properties of specific planar features. Failures often occur along a few distinct planar joints or shears. In such cases, the stability is determined largely by the properties along those surfaces rather than by the average properties of the entire mass. Overbreak, fallouts, and the rock blocks to be supported are strongly dependent on the properties of specific sets of joint and shear zones.

(3) Properties of the intact or laboratory specimens. Some properties of significance in tunneling can be determined from small samples. In many cases, the significant properties can be estimated by combining some of the results from laboratory tests with information on the gross features of the rock mass obtained from exploratory work. The rate of advance of tunnelboring machines is a function of the hardness of an intact sample as welt as the character of the jointing in the rock mass. The compressibility of the rock mass is a function of the modulus of the intact sample as well as the average quality o f the rock mass. The unconfined compressive strength of the sample, when compared with the m a x i m u m in-situ stress, provides an indication of the expected stress-slabbing conditions (rock bursts, popping rock, and formation of new fractures due to high stresses). Some of the slaking and swelling characteristics can also be evaluated from small samples using soil testing procedures. Table I summarizes index properties according to the above categories (Cording et al., 1975). Table II summarizes the index properties which are the basis for the tunnel support classification systems developed by Wickham et al. {1972), Barton et al. (1974} and Bieniawski (1976). In the following paragraphs, several of the index properties related to tunnel support are discussed.

c.

Spacing

Oriented core

f.

Cmtbtnattons of Jotnt sets

Oriented core

Continuity (length)

Evidence of sllckenstdes, cluy. Smoothness of Joints. ([vldence may he limited.)

Log core

weatherl~

e. Attitude (dlp and strlke with respect to tunnel)

d.

gaviness (Inclination. =t" end wavelength)

b. Fllllng end roughness

a

Properties of Malor Joint Sets and Shear Zones

d. Degree of weathering

Seismic ratio Vp fleld.2 ( Vp lab )

Log core: evaluKe Jotnt ~stherlng as well as general

Log core

b.

Frecture frequency (frsctures/foot)

Log core

Core Logging

S. RQO,core recovery

Average RockMassQuahW

Index Prolx~rty

Rock properties of significance in tunneling

TABLE I

Dtroct shear tests of Joint surface or of ft111ng. Determ|nation of p l . t l c l t y Indices.

Evaluate strength, porosity, hardness of sample

Lab sonic velocity (Vp lab )

Lab Index Test

ilorohole CAmerA

6orehole camera

Indtcsttons of soft metertsls fromgeophystca1 loqgtng.

Sonic loqglng (Vp field )

Water pressure tests

In-Hole Tests

Map ortentotlons tn exposures

Map orteetlttons In exposures

Observe extent of Jotnts In outcrops

Observe In exposures

Large scale dtrlct shear tests not usually Wlrrsntnd for tunnels.

Observe tn exposures

Fteld selsmtc (Vp field )

Field Happing, Regtonal Geology, R~ote Sensing, Geophysical, In Sttu Tests

Method of OetemlnaUon

-

Hap ortentettons tn tunnels

limp orientations tn tunnels

Observe extent of Jelnts tn tunnel

Observe tn tunnel

Samptlng and observation of gouge, evtdimce of sltckenstdes, Jolnt surfaces.

Observe tn tunnel

Log m l l of tunnel

Log ~11 of tunw1

Tuneel Exposures

Est 1rote tunnel support requirements

Estlmte field deformtton modulus to evaluate rock dlspluc~ent measurolnents and lining-rock tnteraction Large scale plate loud tests ere sometimes used to evaluate modulus; not usually werrunted for tunnels. Estlmmto tunnel support requtremnts

Application of the Index Property

4.

Slake durability

e.

Plasticity of shales, altered rock or fllllng materials

In-Sltu Permeablhty

g-

f. Boroablllty (for tunnel boring mchlnes)

Swelling properties

d,

Fracturing, open zones In dr1111ng

Presence of hard mlnertls

Deterioration of core

Deterloretlon of

Creep properties

c.

Simple penmb111ty tests for porous mtertsls

Attorperg limits (llqutd, plosttc)

Abrasion hardness tests o lmpect hardness tests, micro-bits

Slake-dureblltty test and plssttctty Indices

Measure swell pressure on samles, evaluate plsstlclty

Constent load trlaxisl creep tests end plastlclty indices

Unconfined cow,prosslon test

Unconfined co~gressive Strength, oc

b.

Evidence of she*red, ~ethered and low qunllty rock

Unconfined c~mpreslon test with strainlmsurements

Properties of the Rock Samp~

S. Nodulus, Eli b

8orohele mter pressure tests, punplng tests

Observe field d r l l l rates

~ --

Evidence of fault zones, civttles, other hlgh permea b i l i t y zouns

Determine extent of rock types

Evaluate extent of she~r zones

Estimate pressures and defomltton In sunl ] tng ground Evaluate ~ n c y for sloktng (time dependent deterioration duo to Imtstdre cMnges)

Obtervst|ons of sunlllng and henve tn tunM1 Observittuns of slektng tn tunnel

Water problems In tunnel

Correlltl With residue1 shelf strength, slaktng, swelling, end creep properties

Evaluate fenstbt 1lt~y of machtne tunneling

Estimate presSUreS and defomttuns on 11ntngs tn squeezing ground

Observations of squeezing

Estimate effect of htgh stresses on s]ebbtng and popptng of rock

Use lab modulus as base for osttmttng the field modulus (see also 1)

--1

118 TABLE II Summary of geotechnical parameters used in tunnel support classification systems Classification system

Parameters

Rock structure rating (RSR) (Wickham et al., 1972)

rock type and weathering general geologic structure average joint spacing orientation of structures relative to tunnel axis ground water volumes and pressures

Geomechanics (RM R) (Bieniawski, 1976)

intact strength RQD joint spacing joint roughness, opening, hardness and filling ground water volumes and pressures orientation of structures relative to tunnel axis

Rock mass quality (Q) (Barton et al., 1974)

RQD number of joint sets roughness of most unfavorable joint sets degree of alteration or filling of most unfavorable joint set stress, intact strength and degree of fracturing ground water volumes and pressures

2.2. Rock quality The average quality o f the r o c k mass is particularly useful in the e x p l o r a t i o n stage for evaluating the c o n d i t i o n s t o be e x p e c t e d along a t u n n e l line. R o c k mass p r o p e r t i e s can be s u m m a r i z e d on a core log b y showing items such as core r e c o v e r y , r o c k q u a l i t y designation (RQD), degree o f weathering, and f r a c t u r e f r e q u e n c y . Such indices m a k e it possible t o rapidly assimilate a large a m o u n t o f e x p l o r a t o r y d a t a and locate zones in the r o c k t h a t are likely t o be o f low quality and p r e s e n t tunneling difficulties. T h e R Q D represents the p e r c e n t a g e o f the NX core r u n t h a t consists o f sound pieces o f core greater t h a n 10 cm (4 inches) in length. (Breaks in the core i n d u c e d by drilling should n o t be considered.) It was originally d e v e l o p e d b y Deere et al. ( 1 9 6 7 ) for evaluating core samples in o r d e r to select the best location for a cavern in a granitic rock. T h e R Q D should be c o n s i d e r e d in the light o f the geologic c o n d i t i o n s and the drilling c o n d i t i o n s at the site. T h e r e m a y be several causes for the low quality o f the core, and t h e y need to be d e t e r m i n e d w h e n using the RQD: (1) Improper drilling and handling. Even t h o u g h t h e drilling breaks are n o t included in the RQD, excessive breakage or loss o f the core during drilling will t e n d to lower the RQD. (2) Drilling parallel to and intersecting a joint. The core t e n d s to break apart w h e n a j o i n t intersects t h e core parallel to the core axis. (3) Separation along bedding planes and foliation surfaces. Even t h o u g h bedding plane partings and foliation surfaces are n o t o p e n in the field, the

119

drilling process and subsequent moisture changes during storage may cause the core to break along partings. Thin-bedded shales c o m m o n l y exhibit this behavior, and if the breaks are included in the estimate, the RQD will be very low. (4) Smaller core size than NX. Additional breakage will take place using smaller diameter core than NX, and the size of the pieces tends to be proportional to the core diameter. (5) Weathering. Weathered zones along joints or in the rock mass will produce lower RQD values. At one site in a granite gneiss, the highly weathered rock could be distinguished from the partially weathered zone on the basis of RQD. The highly weathered zone had RQD values less than 50%, and occurred in the upper 3 m (10 ft) of the rock profile. Below a depth of 6 m (20 ft.), the rock was unweathered, and RQD values were greater than 85%. (6) Shear zones. In unweathered schistose gneisses in Washington, D.C., shear zones or fracture zones were usually present where the RQD for a 5-ft. core run was less than 60%. In some of these cases, slickensides and clay seams were n o t easily observed in the core, but the shear zone could be observed in the exposures in the tunnel at the locations predicted from the low RQD values. (7) Closely jointed and fractured zones. The RQD can be directly related to the degree of fracturing. (8) Core disking. High in-situ stresses may cause disking of the core into thin wafer-like pieces.

2. 3. Properties o f planar features Orientation and strength The stability of a tunnel in loosening ground depends on the orientation of and strength along discontinuities. Some of the more difficult ground conditions can develop where shears and joints are planar and slick, but widely spaced. Thus the RQD in this instance would be high, but the tunneling conditions would be difficult. Orientation of and strength along major continu ous anisotropic features, such as foliation or bedding plane weaknesses, are particularly critical considerations in evaluating rock support conditions. Both the rock quality and the properties of major planes or zones of weakness should be considered in evaluating rock support requirements. Planes that strike within 25 ° of the tunnel axis usually present the most difficult support problem. High-angle joint planes (dips greater than 50 ° ) can cause extensive sidewall support problems. High-angle planes can also form deep, potentially unstable wedges in the crown, particularly if their surfaces are planar and slick, as is c o m m o n to shears and shear zones. Low-angle planes of any orientation and of any strength, will tend to cause shallow slabs to form in the crown of the opening. Minimal support is generally required where there is an absence of well-developed shears, shear zones, or weathered zones, and where there are not enough prominent joint sets to form wedges.

120

(~) DIP ANGLE

0~ . 30 °

H

(AO~ F '

90 °-60"

j (nS)

HEIGHT o f EQUIVALENT ROCK LOAD

(0

"

IS)B

(15 2s)B

30° . 45 °

60 °. 45 °

45 °- 60°

4S°. 30~

( 25 - 43)B

6 0 ° . 7S°

30°- 15"

( 43

M I N I M U M CONDITION FOR FAILURE

Both planes wavy, offset

One p l a n e w a v y or o f f s e t One p l a n e s m o o t h to slightly wavy

One p l a n e sheared, c o n t i n u ous and p l a n a r , One plane s l i g h t l y w a v y

Bath planes sheared, con-

75 ° . 90~

15"- 0°

-

1

>lOB

0)B

tinuous and p l a n a r

L o w l a t e r a l stresses In a r c h ; Surfaces p l a n a r, s m o o t h , poss i b l y open, or progressive f a i l ure a i d e d by s e p a r a t i o n alan low angle joints

Fig.1. Conditions for wedge formation in tunnel crown. The stability of wedges formed in the crown by continuous intersecting joints striking within 25 ° of the tunnel axis is controlled principally by the dip of the planes and the filling on and planarity of their surfaces. Figure 1, based on observations in foliated rocks in New York and Washington, describes the minimum conditions required to form unstable wedges in the crown. The waviness described in the figures is related to the inclination, i, of the large-scale irregularities whose wave length is on the order of 1--10 m (0.3--30 ft.). Wave lengths of this scale will influence the overall stability of both large and small blocks around an opening. The relation between i and waviness is as follows {Cording and Mahar, 1974):

121 i

waviness

00--5 ° 5°--10 ° 10°--20 ° ~20 °

planar slightly wavy wavy very wavy

Failure of wedges formed by low-angle joints (Case 1: 0o--30 ° dip) is possible even if both sides of the wedges are wavy, offset, or rough. It is more c o m m o n , however, if one joint is smooth and planar. For intermediate dip angles (Case 2 : 3 0 ° - - 4 5 ° dip) wedge failures in the crown can occur if one surface is planar and the other offset and wavy. In such a case the wedge will tend to separate from the wavy surface. For joints dipping from 45 ° to 60 ° (Case 3) failure can occur even if one joint is wavy, although commonly, at least one of the surfaces would be sheared. For deep wedges to fail (Case 4 : 6 0 - - 7 5 ° dip), both surfaces must be planar, smooth, and sheared. In this case, both surfaces are at their residual strength. The above discussions of wedge stability in the crown and resulting support requirements assume t h a t displacements are small and that progressive loosening of deeper wedges is not allowed. Loosening in the haunches and sidewalls can also influence the stability of the rock above the crown, particularly in poorer quality rock, such as very blocky and seamy ground where the loosened zone progresses from the walls into the arch along continuous, often planar weaknesses. (For this reason, Terzaghi, 1946, considered both the height and width of the opening in evaluating the height of rock load for steel-supported tunnels.) Planes that strike across the tunnel line will tend to cause support problems at the face, if high-angle joints, and in the crown, if low-angle joints. The orientations shown in Fig.4b are most critical.

Direction of dip with respect to direction of driving The direction of dip with respect to direction of driving affects the stability of wedges in the heading (Cording and Mahar, 1974). Where dips are away from the face (dip direction less than 90 ° to the direction of driving), the plane is first exposed in the arch or crown of the tunnel. Before the plane intersects the tunnel, it is within a few feet of the tunnel surface and may be close enough to form the back of a wedge. On the next round it may be daylighted unexpectedly in the tunnel arch. The wedge backed by this plane will tend to separate and rotate into the heading starting at the face (Fig.2). This may occur unexpectedly after the blast, particularly if the surfaces are planar, smooth and wet. For a plane striking within 25 ° of the tunnel axis, such a condition is most likely to occur on the sidewall or lower arch. It is likely to occur in the crown or arch for a low-angle plane of any strike direction. Planes dipping towards the face (dip direction greater than 90 ° to the direction of driving) are first exposed in the b o t t o m of the face and sidewall, and can be observed as they progress up the wall. A high-angle plane will first

122

FACEAFTERNEXT

i i •RIPTION--r/ DISCONTINUITY/

tDRIVE

PLAN IAL FALL O

5

4

3

2

1

U

~

12

3

45

SUCCESSIVEPOSITIONSOF DISCONTINUITYIN TUNNELFACEASTUNNELIS ADVANCED CROSSSECTIONAT TUNNELHEADING (a) DIP DIRECTIONLESSTHAN 90° FROMDIRECTIONOF DRIVE

(b) DIP DIRECTIONGREATERTHAN 90° FROMDIRECTIONOF DRIVE

Fig.2. Relation of direction of driving to direction of dip. cause a sliding problem on the sidewall, but usually n o t until a substantial portion of the plane is exposed in the b o t t o m of the wall (Fig.2). With further advance of the tunnel, the plane progresses toward the crown and opposite arch. When the plane intersects the arch it may first result in small overbreak. In the next round it is further from the wall and may cause large overbreak in the arch. In the following round, it may be far enough into the opposite wall not to produce overbreak during blasting. This is a critical time, because a wedge may subsequently fall out during the support period. One of the greatest hazards during the support period is the fallouts of large blocks bounded by slick, wet, planar surfaces that provide little or no warning prior to failure.

Combinations o f joint sets to form a wedge In order for a rock block to displace, sufficient joint sets must be presented to form a block of rock that kinematically can move into the opening. One of the more critical support conditions occurs when two sheared, high-angle planes combine to form a deep wedge in the crown of the tunnel. Even though smooth rock joints may bound a given rock block, they may be

123

oriented in such a way t h a t the rock block keys to other blocks and cannot displace into the opening. In this case the rock mass has a strength due to the interlocking effect of the rock blocks. In order for the block to fail it must either shear through irregularities along the joint surfaces, or excavation adjacent to the block must take place to free the block. Minimal support is usually required when there is only one prominent set of joints striking within 25 ° of the tunnel axis. When other intersecting joints are absent, a single shear or joint will tend to cause only shallow slabbing and overbreak near the perimeter of a circular tunnel; large block fallouts are not likely (Fig.3). In a chamber with re-entrants, such as the haunch illustrated in Fig.3, a single joint set can cause blocks to separate over a large portion of the haunch, even in cases where the joint surface is irregular and has a high friction angle. One could conclude from the above example that the design for a large chamber would benefit from detailed exploration to locate all joint sets. This detail is usually more feasible for a large chamber than a long tunnel, because the exploration effort is concentrated in a small area.

2. 4. Stress slabbing Several investigators have used the ratio of the unconfined compressive strength to maximum free field stress, oe/ol, as an indicator of conditions where rock bursts, popping rock, or the formation of new fractures and slabs develop around an opening (Cording et al., 1971; Cook, 1973; Barton et al., 1974; Bieniawski, 1976). Stress conditions required to cause minor bursting in chambers in crystalline massive rocks in the walls of Norwegian valleys, were described by SelmerOlsen (Brekke, 1970) in terms of the tensile strength of the rock and the vertical distance between the chamber and the top of the valley side. A ratio of tensile strength, Or, to in-situ stress can be obtained by expressing the relationship in terms of an equivalent overburden stress, or, due to a column

\ 50'

24'

Fig.3. Influence o f single joint set on overbreak.

124 of material having a height equal to the distance to the top of the valley side. For ot/a,, less than approximately 0.5, bursting would be expected (equivalent to a oc/o v less than 10, assuming oc is 20 times or). Although stress conditions near this limit do not produce major rock bursts, they may be of substantial economic consequence in the excellent quality rock because of the requirement for support -- even light support -- in a cavern that would otherwise be unsupported (Brekke, 1970). Other investigators indicate that rock bursts or stress slabbing may develop for values o f a c/ol less than approximately 4 to 6 (Cording et al., 1971; Cook, 1973; Barton et ah, 1974). In high strength quartzites, a¢ = 200 MPa (30,000 psi), Cook (1973) indicates that a ratio of o~/ol less than 1.5 results in inability to keep a mine open. In a low strength rock, such as a t u f f having % = 10 MPa (1500 psi), the same ratio would produce new fractures and slabs around the openings as they are excavated, but a much smaller release of energy. Popping rock rather than major bursts would be expected (Cording et ah, 1971). In heavily jointed and sheared rock, deformations take place along the existing discontinuities, and rock slabbing or bursting is not as evident. However, massive rock adjacent to fault zones and fractured rock may exhibit prominent stress-slabbing conditions. Some of the early work regarding stress concentrations around openings placed emphasis on determining maximum stress concentrations and shaping the opening to reduce the stress concentrations to a minimum. However, the location of the highest stress concentration occurs around the smallest radius curves and does not involve a large volume of rock. Of more importance are the zones where large stress differences exist to depth, behind the large radius surfaces, such as the high side walls of a powerhouse. The stability of such areas can be even more critical when joints or bedding planes parallel the surface. For example, although stress slabbing developed to a depth of 1--2.5 m (3--8 ft.) on the surfaces of a 30-m (100-ft.)-diameter rock bolted cavern in a highly stressed t u f f (~c/ol = 1.5) the most critical condition occurred on the 36-m (120-ft.)-high planar sidewall where a high-angle joint set almost parallel to the wall combined with new fractures due to high stress to allow large block movements to a depth of 6--9 m (20--30 ft.) (Cording et al., 1971). Nearby, in an identically shaped second chamber with the same initial rock-bolt support, joints did not parallel the wall and stability problems did not develop. Conversely, had the cavern been constructed in a rock mass with the same joints almost parallel to the wall, but in a rock of higher intact strength (for example oclol > 6), it is doubtful that major stability problems would have developed. From their experience in caverns in Norway, Selmer-Olsen and Broch (1977) emphasize the importance of minimizing the length of planar walls of a cavern that are parallel to the direction of the maximum principal compressive stress. They show desirable cavern configurations to minimize the support requirements for both highly stressed and moderately stressed ground, for various ratios of horizontal to vertical stress.

125 In shallow chambers, and even in many chambers at depth, the stress levels are low enough that major difficulties with stress-slabbing do not develop. In these cases, the opening shape and the required rock support are primarily determined on the basis of the geometry of the rock discontinuities so that the critical rock wedges (such as those described in Section 2.3) are properly supported. 3. SUPPORT

3.1. Summary of support selection The selection of a support involves choosing a system compatible with the ground conditions, excavation procedures and sequences, and with the materials, equipment, and labor force available to the contractor. Indices such as those described in Section 2 are useful for defining the ground conditions so that appropriate support is chosen. The following paragraphs describe the factors involved in selecting the support system. They are: (1) Heading support. The support must be capable of being installed safely and providing adequate support in the heading. (2) Support spacing. The support must be spaced closely enough to prevent excessive loosening and fallout from between the support elements. (3) Support displacement. The support system should be capable of limiting rock movements to acceptable levels and deforming under the anticipated rock movements without failure. (4) Support pressures and bolt lengths. The support system should have sufficient capacity to hold the critical rock wedges in place and rock bolts should be of sufficient length to develop anchorage behind the critical wedges. (5) Compatibility of initial and final support and integration of excavation and support sequences in large chambers. The initial and final supports should be coordinated so that they are compatible. In some cases, both the initial and final supports can be provided by a single system. However, there are cases where a single system cannot satisfy the functions of both the initial support (stabilization of the heading) and the final support (support the tunnel for its intended life), without excessive deterioration. In large chambers, initial excavation sequences may be used as access areas to help support later excavation stages.

3.2. Heading support In underground construction, the support must be integrated with the excavation sequence. Support size and spacing is determined not only by the anticipated rock loads, but also by the requirements for protecting workmen by providing timely support in the heading.

126

The reach, or maximum distance excavated ahead of the last support (Fig.4a), is a critical factor in the stability of a heading. The reach can be minimized by placing support at the face, and excavating only short distances in each round. Specifications for support should not only include the spacing and capacity of the support system b u t the maximum distance from the face that it is to be installed. The height and width of a tunnel determines the distance ahead of the face that a failure can occur. To minimize the possibility of failure ahead of the face, the face height can be limited by excavating a smaller tunnel, by leaving a core in the center of the excavation, or by sloping the face. Spiling can also be used to provide support ahead of the face and to support the crown before it is exposed by excavation (Fig.4b). In soft ground, spiles can be driven ahead of the face. In rock, spiles, consisting of reinforcing bars installed in percussion drilled holes and fully encapsulated in resin have proven effective in providing support at and ahead of the face.

i~Reach.i r 7 Steel Rtbs

~,

!

~aJ~ ! ~

Reach

I

I I

,, I

__J

~j

........ ..Q .....

Fully grouteds

p

/

~

l

/ 7/¢ Fig.4. Support at heading, a. Examples of reach, b. Use of spiling for support ahead of the face.

127

3.3. Support spacing Spacings for rock bolts typically range from 1 to 3 m (3--10 ft.), with the wider spacings used in excellent quality rock or in rock having an intermediate support system such as shotcrete and mesh. Rock-bolt spacings should also be less than 1/2 the bolt length (Lang, 1962; Dept. of the Army, 1975). In most small tunnels, and in some large chambers in excellent quality rock, if the selected supports are spaced so that they provide adequate protection, they usually have sufficient capacity, or can be easily dimensioned, to carry the anticipated rock loads. This is not the case in very large chambers or in intermediate-sized chambers in poor rock.

3.4. Support displacement Most rock tunnels and chambers, particularly those at shallow depth, are located in loosening ground. For such tunnels, the support should be placed as soon as possible to minimize rock loads. The loads that develop will be a function of the geometry and strength of the discontinuities, the displacements allowed, and the size and geometry of the opening. Ground reaction curves for elasto-plastic or creep-sensitive materials, such as those described by Rabcewicz (1969}, show that the pressure required to support the opening decreases as the inward displacement is allowed to increase (Fig.5a}. Delay in placing supports results in less load on the support. In many loosening ground conditions, however, the rock mass is stiff enough so that only small displacements are required to reduce the pressures to minim u m levels, and the support placement need not be delayed to develop minimum pressures. In fact, any delay in placing the support will result in loosening of the ground and a reduction in the strength of the mass, so that the rock loads increase (Fig.5b). In shallow rock chambers, it is particularly important that loosening be limited by early placement of support so that the thin rock arch is preserved.

3.5. Estimates o f support pressure For a frictional rock material loaded by its o w n b o d y forces, the support pressure, P,, in the arch of a tunnel will be:

P~ = nB~ where B is the width of the opening, ~/the unit weight of material, and n is an inverse function of the shear strength of the rock mass and its discontinuities. The shear strength may increase or decrease with opening width. If it is assumed that the shear strength does not vary with width, then an increase in width requires a proportionate increase in the support pressure. Several factors are described in Section 3.7 that influence the relation between n and opening width. The value of n also increases when loosening and displacements

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Support Pressure, Pi

Displacement

Support \ Pi ~

Pressure,

Support InstalIe

d

Ol

~

, .

b

DIsplacenmnt

d

Fig.5. Ground reaction curves, a. Squeezing ground, b. Loosening ground.

cause a loss of interlocking and confinement of rock blocks or a reduction from peak to residual values along joint surfaces bounding the blocks. The height of rock loads, Hp, for steel supports in loosening ground is given by Terzaghi (1946} as a function of the width, B, and height, H t, of the tunnel. In good quality rock, Hp ranges from 0 to 0.5B whereas in poorer quality rock (blocky and seamy to crushed ground) Hp ranges from 0.25B to 1.1(B + Ht), equivalent to n values ranging from 0.25 to 2.2. Large rock-bolted chambers in fair to excellent quality rock typically have n values in the arch ranging from 0.1 to 0.3 (Cording et al., 1971). Rock-bolt lengths typically range from 1/4B to 1/3B. Rock loads will be smaller for support systems that limit the ground movements, but the height of rock load (or rock pressure} should remain a function of the width of the opening. In large rock chambers, rock-bolt lengths and pressures with respect to opening height (on the sidewall} or width (in the arch) are typically lower in the

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sidewall than in the arch. However, the high sidewalls used in many large chambers can present a more critical stability condition than the arch when major shear zones or joints intersect the chamber surface. Substantially longer bolts and higher support pressures may be required on the walls when adversely oriented joints or weak zones are present. Since Terzaghi's work, many rock-classification systems have been developed for use in evaluating tunnel ground conditions and support requirements. A large number of classifications are developed for specific sites, and are appropriate for specific geologic and construction conditions. Such classifications provide a basis for communicating between various parties engaged in the project. Other investigators have developed systems for more general use. Lauffer's well-known classification (1958) relates standup time to tunnel rock behavior and opening width. More recently Linder (1963) has correlated Lauffer's rock descriptions with support design for use in tunnels. Deere et al. (1969} summarize typical support types and dimensions for 6--12 m (30-40 ft.)-wide tunnels as a function of the RQD. Wickham et al. {1972), Barton et al. (1974) and Bieniawski (1976) have each developed classification systems that include RQD and other significant rock indices. From these indices a classification number can be obtained that is correlated with support requirements. The description of significant index properties is one of the most useful aspects of the classifications. In many cases, the individual indices may be of more benefit in describing potential tunnel ground conditions than a combined classification number, particularly when using the systems with different geologic and construction conditions.

3.6. Approach to selection o f support Although classification systems can be used to obtain an approximate estimate of support requirements, the procedures for evaluating support requirements should include the following steps: (1) Evaluate geology and determine significant index properties. (2) Outline expected ground behavior and its effect on construction. This additional step requires a knowledge of construction procedures as well as the models of rock behavior that are appropriate to the given geologic setting and excavation geometry. The geologic data should be plotted in plan and crosssection views of the proposed opening for each excavation sequence in order to identify potential problems and to check the adequacy of the proposed support systems. (3) Select support systems and construction procedures that permit the headings to be safely and economically excavated, and that provide permanent support suitable for the intended life of the facility. In loosening ground, support requirements can be estimated by evaluating the critical rock wedges (such as those illustrated in Fig.l) and designing the support to hold the wedges in place. Exploration should be carried out to determine the geologic features that will affect wedge stability. For large openings, exploration is concentrated in

130 a relatively small area. Not only should the average quality of the rock throughout a large chamber be determined, the major shear zones and other low-quality zones should be specifically located and oriented with respect to the walls and intersections of the chamber. Support of benches, haunches, and intersections in the opening may be affected by a single, unfavorably oriented joint set and consequently the orientation of all major joint sets should be determined in the exploration. Although information can be obtained from borings, the most complete exploration information is obtained from a pilot tunnel driven at the location of the chamber. In long tunnels it is not usually possible or necessary to locate all major shear zones. An estimate should be made of the relative percentages of the various tunnel ground conditions expected. The orientation of the major structures and an estimate of their frequency and character should be determined for given reaches of the tunnel. Once the geometry and strength of the rock mass and its discontinuities have been estimated, critical wedges can be selected on the basis of criteria such as those outlined in Fig.1. Analyses of displacements and stress around the opening, using finite element models with joints, may also provide insight into the extent of the critical wedges that must be supported in highly stressed ground. Proposed excavation and support sequences should be tested against the given wedge geometries to determine if the chamber will remain stable as it is excavated. Support pressures are selected so that the stability of the critical wedges is maintained.

3. 7. Size effect In the writers' experience, large chambers have required substantially greater support pressures than smaller openings, and the support pressures have been related to the size of the opening. For example, in the cavern in tuff (oc/ol = 1.5, RQD: excellent), a 0.14 MPa (20 psi) pressure was required to stabilize the 36-m (120-ft)-high sidewall while typical rock-bolt spacings and pressures [0--0.03 MPa (0--5 psi)] were adequate on 3--6-m (10--30 ft.)high walls. In Washington, D.C., in schists with planar and continuous joints and shears, rock pressures in the arch of 21 m (70-ft.)-wide openings measured 0.14 MPa (20 psi), while rock support in the r o o f of 6-m (20-ft.)-wide openings provided a support pressure of approximately 0.03 MPa (5 psi). Barton et al. (1974) compare support pressures for openings of varying size with a combined rock mass quality parameter, Q, but do not include the effect of size on support pressure. In cases where the spacing of the support controls, the support pressures may not appear to be a function of width of opening. For example a 6 m (20-ft.)-tunnel and a 12 m (40-ft.)-tunnel may both be supported with 25 mm (1-inch)-diameter rock bolts on a 1.5 × 1.5 m (5 ft. × 5 ft.} spacing. The spacing is controlled b y the quality of the rock and the potential for fallout between bolts. Even larger chambers, if in excellent quality rock where only local loosening takes place around the perimeter, may require small support pressures and be adequately supported by the

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above bolt pattern. This may be the case for the Norwegian chambers described by Brekke (1970}, where rock support, if used at all, is nominal and bolt pressures required are small. Some of the factors that affect the relation between support pressure and opening size in loosening ground are outlined below and illustrated in Fig.6: (1) Joint spacing with respect to size o f opening. For small openings with widely spaced joints, the rock mass acts as if it has a cohesive c o m p o n e n t of strength, and the pressure, P1, required to support the opening is reduced by a cohesive term, C:

Pi = nB~--C where C is a function of the ratio of joint spacing of the critical joint sets to the tunnel diameter. In a small opening, little or no support may be required,

a

c

Fig.6. Geologic features affecting size. a. Influence of joint spacing, b. Influence of excavation increment. Influence of joint continuity.

132 while in a large opening, where joints are closely spaced with respect to the width of the tunnel, the cohesive component approaches zero, and support pressures are large. {2) If a large chamber is excavated and supported in increments, the loosening of the rock around the opening will primarily be a function of the width of the individual excavation increments, whereas a small tunnel excavated full face will have a loosened zone proportional to its full width. However, if through-going discontinuities are present in both cases, the rock pressures may still be proportional to the width of the opening. (3) If the continuity of shears or joints is substantially less than the width of the opening, then large wedges may not be able to form, and the rock pressures will not increase in direct proportion to the width of the opening. The wavelength of the irregularity with respect to the width of the tunnel is another scale effect that influences pressure. In the Washington, D.C. chambers, the joints and shears are planar and continuous so that the pressures tend to scale with the width of the opening. (4) Support of a large tunnel is made more difficult not only because of increases in rock pressure, and changes in the scale of the joint spacing with respect to opening width, but also because of the increased construction problems in excavating and installing support, and the limitations of the support system. For example, if a straight-leg steel support is placed in a small tunnel, it can be braced to provide resistance to side loading by placing an invert strut across the b o t t o m of the tunnel. In a tunnel with a high sidewall, a heavy straight-leg steel support would not be capable of providing substantial lateral support even when an invert strut is installed. In a large tunnel heading, not only are the rock loads increased, but the size of the face and the disturbed zone ahead of the face is enlarged, making it more difficult to install the support safely in the heading.

3.8. Excavation and support sequences in large chambers In a large chamber, excavation and support sequences are more complex than in a small tunnel, and have a greater influence on the permanent support requirements. Support procedures should be planned to provide maximum support prior to excavation of the main portion of the chamber, and to provide initial and permanent linings that are compatible, and do not duplicate support functions. Excavation and support sequences for large chambers excavated at shallow depth are particularly critical, because of the need to maintain the thin rock arch and prevent collapse to the ground surface. Large shallow chambers constructed before the advent of rock bolts, such as the subway stations built in New York City at the turn of the century, were supported internally with steel or timber, and were supported, and resupported with posts and beams as the various excavation stages were carried out. With rock bolts, the internal supports can be minimized, and portions of the rock arch can be pre-supported prior to excavation. Figure 7 illustrates the construction sequence for a subway station in rock, similar to

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the procedures used for the 18-m (60-ft.)-wide stations in the foliated schists on the Washington, D.C. Metro, where rock cover over the crown may be as little as 9 m (30 ft.). The running tunnels (R) are excavated with tunnel-boring machines through the station area prior to the construction contract for the main station. The running tunnels and pilot tunnel (P) provide an opportunity for rock-bolt support to be placed, n o t only immediately adjacent to the existing excavations, b u t also over the arch of the station, above the areas y e t to be excavated (Figs.7a and 7b). The bolts perform a similar function as the spiling in Fig.4, b u t are placed laterally over the arch prior to excavating the main portion of the station. Once the rock arch is pre-supported with fully grouted rock bolts, the upper heading can be excavated, placing the arch support on the foundations set in the running tunnels (Fig.7c). The height of the upper heading and volume of material excavated in this stage is greater than would be possible without the bolted arch, because the grouted bolts in the arch have reduced the potential for failure above and ahead of the face. In one of the station excavations in Washington, it was observed that the face of the excavation had broken ahead approximately 3--4.5 m (10--15 ft.) further than the ends of the blast holes. The fallout of rock was confined to the excavation limits by the rock bolts, and did not proceed above the arch, as would have occurred had the bolts n o t been present (Cording et al., 1977). Because of the shallow cover and the presence of steeply dipping shear zones striking parallel to the station axis, the rock bolts alone do not provide sufficient support for the arch. A final lining of light steel ribs encased in shotcrete or concrete is installed as the upper heading is excavated. The amount of support required using this procedure is much less than would be required if heavy steel ribs were installed to support the arch, and then the permanent concrete lining were placed at a later date, after excavation was completed. 4. OBSERVATIONS

4.1. Purpose and method o f observations The observation program for a large chamber differs from the program for a tunnel. A large chamber opened b y excavations in sequence is much more difficult for the miners and engineers to observe, compared to a small tunnel that is opened full face and immediately supported. Thus, more reliance must be placed on measurements of rock movement and support performance with a large chamber. R o c k displacement measurements can provide early warning of potentially unstable conditions. For a large chamber, opened in increments, there is an. opportunity to monitor the displacements during several sequences of excavation. Movements and support loads observed during one sequence can be used to determine the adequacy of method for the next sequence, and corrections to the excavation and support procedures can be made, if needed.

b~

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Chambers are sometimes designed to take advantage of observations by installing a minimum support, then adding more support where observations indicate it is needed. In most civil-works chambers in the U.S., it is preferable to design an adequate support system, in which minimal construction changes will be required, and then use the observations to confirm design assumptions, or to warn of any abnormal conditions that may require adjustments in excavation or support procedures. Instrumentation should be considered where openings are large, where there is a difficult and complex geology, where the opening geometry and the required construction procedures are complex, where there is a possibility of large movements that could cause stability problems or damage to adjacent structures, or where information is needed on support loads to aid in designing future sections. Observations in a rock chamber or tunnel during construction also provide a means for relating index properties and other significant geologic and construction parameters to the performance of the chamber. An instrumentation or observation program is of little benefit to the profession unless the results are compared with the geologic index properties at the site. In many cases, data originally t h o u g h t to be variable and scattered fall into place when the details of the geology are understood. In many cases, an understanding of the performance of a tunnel can be gained w i t h o u t the use of extensive instrumentation; but to do this careful visual observations of such occurrences as overbreak, fallout, rock loosening, fresh fractures, cracks in the lining, and other evidence of rock movement and support distress should be made and compared with the behavior of instrumented sections. One of the most informative and reliable measurements in a tunnel or chamber in rock is the measurement of displacements using extensometers. Displacement measurements integrate the local strains and displacements over the measurement length and thus do n o t present the same difficulties as strain gages or pressure cells, which measure the conditions at a point and which tend to produce erratic and scattered results. The displacement measurements give an indication of the stability of an opening. Impending failures or fallouts are directly signalled by measuring those rock displacements that would go unrecognized by the naked eye. Sufficient data has been collected on rock movements, and correlations with analytical models are available, so that observed displacement measurements can be interpreted. The anchors for extensometers can be placed in such a way that information can be obtained as to the depth of the movement as well as the magnitude. The planning of a remedial support system is greatly benefited by having information on the volume of rock which is involved in the movements Fig.7. Excavation and support sequence, a. Excavation of running tunnels (R) and pilot tunnel (P) and installation of rock bolts, b. Enlargement of pilot tunnel (Stage 1 ) excavation of corner (Stage 2) and installation of wall plate (Stage 3). c. Excavation of main heading (Stage 4). Installation of arch (Stage 5). Removal of remainder of invert (Stage 6).

136

Extensometers should be installed before significant rock movements take place. The extensometers installed in boreholes from the ground surface or from existing nearby excavations are best suited for this purpose, as they can be installed prior to any excavation in the vicinity of the instruments. Extensometers can also be installed in small pilot tunnels located in the crown of the proposed excavation. Subsequent movements due to driving the larger openings are then fully recorded as the heading approaches, reaches, then passes the instrument location. If the above alternatives are not possible, then extensometers can be installed from within the tunnel near the face of an advancing excavation. Obviously, displacements occurring prior to installation of the extensometer -- in front of the heading and as the heading is first opened -- are not recorded. Such extensometers are usually installed from the same work platform that the contractor uses to drill rock-bolt or blast holes, and they can be installed near the heading after the initial support is placed. Displacements during the first round of advance of the heading away from the section can then be recorded, if the extensometer is of the type that can be installed rapidly and read immediately after installation. Extensometers are available that can be installed and read within 30 minutes after the borehole has been drilled. Most extensometers can be read with a repeatability of +0.025--0.05 mm (0.001 to +0.002 inches), which is quite adequate for evaluating the significant rock movements around a tunnel or chamber. After the geologic observations and the rock-displacement measurements are organized, the next most important consideration is the performance of the support itself. Instruments are available to measure distortion of a lining, strain in the section of a lining or the total load on the lining. Much useful information has also been observed by noting cracks in shotcrete, loosening of rock blocks, and the distortion of the supports. Again, the performance of supports may be variable and depend on local geology, which should be mapped in detail.

4.2. Criteria for evaluating displacements Criteria for determining if displacements are indicative of either stable or potentially unstable behavior are outlined in this section. Displacements that indicate local instability, such as loosening of a thin rock slab, are distinguished from those that indicate a more widespread and deep-seated condition affecting the stability of the entire opening. In most cases, a single criterion is not adequate for evaluating the stability of a tunnel. Displacement measurements are most valuable when extensometers are installed at or before the beginning of excavation, and when measurements have been taken regularly throughout the entire excavation period at several locations so that a complete history of movements is available. They will be of use if the geologic conditions and construction events in the vicinity of the measurements are also recorded and compared with the movements.

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Displacement magnitude Elastic or elasto-plastic continuum solutions are quite useful for comparison with the observed displacements in a rock tunnel or chamber, even though the rock mass may suffer large displacement along joint surfaces and not behave as a continuum. The continuum solution is valuable because it provides an estimate of the displacements the mass would undergo if loosening along the joints were minimized. Unstable conditions may exist if displace ments are large with respect to the displacements predicted from elastic theory. Either a closed elastic solution, assuming simple boundary conditions, or a finite-element elastic solution that approximates the more complex boundary conditions in a chamber, can be used to estimate the elastic displacements. The rock movements measured in deep tunnels and chambers, such as those constructed for powerhouses, are typically on the order of 1--3 times the movements one would calculate from elastic theory, using an appropriate in-situ rock modulus that accounts for the stiffness of the joints in the rock mass (Deere et ah, 1967). Normal displacements in most of these chambers were in the range of 2.5--7.5 m m (0.1--0.3 inch). Movements that occurred where shear zones or other major discontinuities were not adequately supported were typically in the range 12.5--75 mm {0.5--3.0 inch), approximately 5--10 times the elastic displacements. In most cases, movements of this magnitude were recognized as being excessive, and corrective measures were taken (Cording et al., 1971). Displacement measurements previously obtained in stable sections of a tunnel can be used to establish the typical behavior to be expected in other portions of the project. For example, in the shallow Washington Metro tunnels where the in-situ stresses were low, the calculated elastic displacements were extremely small -- on the order of 0.25--1.25 mm (0.01--0.05 inch). Observed displacements for well-supported excavations were typical, 2--5 times the elastic displacements, and ranged from 0.5 mm (0.02 inch} for small tunnels to 5 mm (0.2 inch) for larger excavations. Displacements of 0.5--1.25 mm (0.02--0.05 inch) were typical for 1.2--1.8-m (4--6-ft.) wide rock blocks supported by a 50-mm (2-inch) thick shotcrete layer and fully grouted, non-tensioned rock bolts (Cording et al., 1977). The displacements represented the small a m o u n t of separation along the joints that occurred as the tension increased in the bolts. The movements would not have been substantially less than 0.5--1.25 mm (0.02--0.05 inch), even if the calculated elastic displacement had been much smaller.

Rate of displacement Displacement rates should be examined closely when evaluating the stability of a tunnel. Sudden increases in the rate of rock movement that are larger than would be expected for the increment of excavation carried out in the vicinity of the extensometer may provide an early indication of an

138

unstable condition. High rates of movement that are unrelated to excavation or that continue after the face has advanced well beyond the extensometer location, may also indicate an unstable condition. One of the best means of evaluating such rates of movement is to compare them with rates previously observed in portions of the tunnel that were well supported and where the displacements ultimately stopped. In a shallow subway station in Washington, displacement rates in wello supported portions of the cavern were less than 0.1 mm {0.004 inch)/day when the heading had advanced one tunnel width, and less than 0.025 mm (0.001 inch)/day after the heading had advanced three or more tunnel widths. Rates of movement that were 3--4 times these values were usually indicative of conditions where additional support was required (Cording et al., 1977). Extensometers with anchors at several depths provide useful information that a single position extensometer or a convergence gage cannot provide regarding the depth of the movement zone and the volume of rock involved in the movements. Loosening of shallow slabs is c o m m o n as an excavation is opened, but if the loosening is allowed to continue, the zone of movement will tend to extend to greater depths, and the rock loads may increase. Extensometers should be long enough to extend beyond the potential zone of movement, or supplementary measurements, such as precise surveys, should be made. In the underground Machine Hall at Morrow Point Dam, 50.8 m m (2 inches) of displacement took place on one wall of the chamber. The wedge was so large that a 15-m (50-ft.) long extensometer on the wall registered no displacement; it was located entirely within the moving wedge. The movements were detected by precise survey measurements in the chamber (Dodd, 1967).

Displacement capacity of the support system Observed displacements should not exceed the displacements that will cause distress or failure of the support system. Shotcrete has been observed to crack in the tunnels of the Washington, D.C. Metro when the differential movement between rock blocks exceeds 1.2--2.5 mm (0.05--0.10 inch). At the Nevada Test Site, bearing plates dished and eleven rock bolts broke when rock displacements approached 50 mm (2 inches). (The bolts were tensioned, but not grouted, over an 8-m (24-ft.) length.)

Displacement capacity of the rock mass Displacements should not exceed the capacity of the rock mass to maintain its strength and coherence, unless the support system is capable of supporting the resulting increased rock loads. Rock strength along joints decreases with displacement as irregularities on the joint surface are sheared or overridden. In the Washington Metro Subway chambers, relatively small displacements at the edge of the excavation caused movement of large volumes of rock. In one case, in a 6-m (18-ft.)-wide by 8-m (24-ft.)-high tunnel, large sidewall loads developed at displacements of 2.5--12 mm (0.1--0.5 inch) because the discontinuities were oriented unfavorably (striking almost parallel to the wall and

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dipping toward the excavation) and were planar and continuous. The sidewall movements also caused loosening of rock to a distance of 6 m (20 ft.) above the crown. The magnitude of the displacements that will cause loss of strength of the rock mass depends on the a m o u n t of displacement it takes to override or shear off the irregularities so that the block will fall out or so that residual strength will be reached on the joint surfaces. Where joints are planar and slickensided, and sufficient joint sets are present to form blocky rock, the displacements required to cause failure will be less than those where surfaces are irregular and joints are discontinuous. This range of displacements is estimated to be approximately 2.5--25 m m (0.1--1 inch) in Washington, D.C.

4. 3. Supplemental observations Supplemental observations of support load, support distress, and rock movement will aid the interpretation of displacement measurements. Some of these observations are outlined below: (1) Opening of joints or movement of rock blocks. Such observations can be made visually or with a tape measure or survey. Open cracks in boreholes can be inspected with a stratascope. Rock displacements are also indicated by offsets in open boreholes drilled prior to the time of the rock movements. Any loose slabs in the vicinity of an extensometer should be noted. (2) Mapping of joints, shear zones, and other geologic features that could cause movement. Observations of overbreak and rock loosening along the joints and shear zones will aid in evaluating the significance of these features. (3) Crack surveys in shotcrete. The width, length, and relative movement of the crack should be measured with time, and the thickness of the shotcrete in the vicinity of the crack determined. (4) Evidence of distress or displacement of steel ribs and timber blocking. Crushing, bending, or loosening of the timber should be noted. Distortion and twisting of rib sections and opening of b u t t plates can be measured with a tape. Deflection or settlement of the ribs can be measured by survey. Closure of ribs can be measured with a tape extensometer or a tape measure. (5) Evidence of distress or loosening of rock bolts. Tensioned, non-grouted bolts should be checked for loosening of the bolts at the bearing plate, bending of the bolt head, or breaking of the bolt. Load cells installed on nongrouted bolts can be used to evaluate the adequacy of the bolt installation, but provide relatively little information on the overall performance of the rock bolts. Grouted bolts are usually preferred to non-grouted bolts, and they cannot be instrumented with load cells. (6) Measured strains or loads on the support system. Strain gages can be attached to steel ribs or embedded in shotcrete or concrete. Load cells can be placed beneath posts. Gages that will be stable in the tunnel environment should be selected. (7) Groundwater flows and pressures. The location and quantity of flow from rock joints or from cracks or joints in the lining should be noted.

140 Piezometers can be installed in the rock or behind the lining to measure water pressures. 6. CONCLUSIONS The in-situ methods emphasized in this paper are used to relate rock index properties to the performance of a tunnel or large chamber. The methods are carried o u t by exploration prior to construction and b y observations during excavation. In most cases, the estimation of tunnel stability and support requirements is benefited more by careful observations of the geology and its influence on tunneling than b y an extensive program of in-situ testing. Index properties of significance in tunneling can be divided into the following categories: the average quality of the rock mass, the properties of planar discontinuities, and the properties of intact specimens. The average quality of the rock mass can be determined from quantitative estimates of rock quality, fracturing, and weathering in the rock core. Calibrated descriptive estimates may suffice. The average rock quality indices are particularly useful in the exploration stage in locating the low quality zones that may cause tunneling difficulties. Properties of planar discontinuities, such as the thickness of clay filling, may be lost unless procedures for obtaining full core recovery are carried out. The large-scale features of planar discontinuities, such as continuity and waviness can only be observed in exposures. Pilot tunnels not only provide exposures of the geology but can also serve as an in-situ test in which the influence of rock properties on overbreak and support requirements can be studied and used as an indication of conditions to be expected in a large chamber. However, pilot tunnel support requirements may not be related directly to those for a large tunnel because of size effects. The amount and type of support required may differ considerably. As described in Section 3, support pressures in loosening ground will tend to increase in proportion to the width of the opening. The relation is affected by the geologic conditions illustrated in Fig.6; joint spacing with respect to opening width, the continuity and waviness of joints, the size of the individual excavation sequences, and the timeliness of support placement. In comparing a small tunnel with a large tunnel, it should be recognized that not only will support pressures differ b u t there will be a significant difference in the construction methods required to stabilize the heading and install supports. Extrapolation of the observations from a small to a large tunnel can be most appropriately made by "taking apart" and examining the geologic features and construction procedures in the small tunnel, then putting these features back together in the larger tunnel. The effect of the geologic indices on the performance should then be examined by proceeding through potential construction sequences for the large tunnel, step by step. One may find, in "taking apart" and then putting the elements back together that pieces are missing or do not fit. Such an exercise will clearly highlight the areas in which behavior or support conditions are unknown.

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A similar approach can be taken when using core borings and other exploration data to estimate support and construction requirements. Although classification systems and rules of thumb that directly relate rock indices to support requirements can be useful, it is the writers' conclusion that an additional step is needed in the evaluation of support: (1) Evaluate geology and determine significant index properties. (2) Outline expected ground behavior and its effect on construction. This additional step requires a knowledge of construction procedures as well as the models of rock behavior that are appropriate to the given geologic setting and excavation geometry. The geologic data should be plotted in plan and crosssection views of the proposed opening for each excavation sequence in order to identify potential problems and to check the adequacy of the proposed support systems. (3) Select support systems and construction procedures that permit the headings to be safely and economically excavated, and that provide permanent support suitable over the intended life of the facility. Much can be gained by observing overbreak, fallout, evidence of slabbing due to stress, rock movement, and support distress. The information should be collected at the heading of the tunnel, as the rock is excavated and supported. Observations of support procedures immediately after blasting are particularly important. Rugged, reliable instruments that can be installed, read, interpreted, and related to geologic conditions in a timely manner can contribute much to the observation program. REFERENCES Barton, N., Lien, R. and Lund, J., 1974. Engineering classification of rock masses for the design of tunnel support. Rock Mech., 6(4): 189--236. Bieniawski, Z.T., 1976. Rock mass classifications in rock engineering. In: Z.T. Bieniawski (Editor), Exploration for Rock Engineering. Balkema, Rotterdam, 1: 97--106. Brekke, T.L., 1970. A survey of large permanent underground openings in Norway. Proc. Int. Syrup. Large Permanent Underground Openings, Universitetsforlaget, Oslo, pp. 15--30. Cook, N.G.W., 1973. The siting of mine tunnels. Assoc. Mine Managers, No. 3/73. Cording, E.J. and Mahar, J.W., 1974. The effects of natural geologic discontinuities on behavior of rock in tunnels. Proc. Rapid Excavation and Tunneling Conf. AIME, San Francisco, 1: 107--138. Cording, E.J., Hendron Jr., A.J. and Deere, D.U., 1971. Rock engineering for underground caverns. Syrup. Underground Rock Chambers, Phoenix, ASCE, pp.567--600. Cording, E.J., Hendron Jr., A.J., MacPherson, H.H., Hansmire, W.H., Jones, R.A., Mahar, J.W. and O'Rourke, T.D., 1975. Methods for Geotechnical Observations and Instrumentation in Tunneling. NSF Rep., UILU-ENG 75 2022, University of Illinois, Urbana, Ill.,Vols. 1 and 2. Cording, E.J., Mahar, J.W. and Brierley, G.S., 1977. Observations for shallow chambers in rock. Int. Syrup. Field Measurements in Rock Mechanics, Zurich, pp.485--508. Deere, D.U., Hendron Jr.,A.J., Patton, F.D. and Cording, E.J., 1967. Design of surface and near-surface construction in rock. Syrup. Rock Mech., 8th, Minnesota. Deere, D.U., Peck, R.B., Monsees, J.E. and Schmidt, B., 1969. Design of Tunnel Liners and Support Systems. U.S. Department of Transportation, Washington, D.C., Contr. No. 3-0152.

142 Department of the Army, Corps of Engineers, 1975. Rock Reinforcement in Civil Engineering Works. Rep. EMl110-1-2907. Dodd, J.S., 1967. Morrow Point Underground Powerplant Rock Mechanics Investigations. A Water Resources Technical Publication. U.S. Department of the Interior, Bureau of Reclamation, Denver, Colo. Lang, T.A., 1962. Theory and practice of rock bolting. Trans. AIME, 23. Lauffer, H., 1958. Gebirgsklassifizierung fiir den Stollenbau. Geol. Bauwes., 24: 46--51. Linder, R., 1963. Spritzbeton in Felshohlraumbau. Bautechnik, 40(10): 327--331. Peck, R.B., Hanson, W. and Thornburn, T.H., 1953. Foundation Engineering. Wiley, New York, N.Y., 410 pp. Rabcewicz, L.V., 1969. Stability of tunnels under rock load. Water Power, 1969 (June-August). Selmer-Olsen, R. and Broch, E., 1977. General Design Procedure for Underground Openings in Norway. Rockstore 77, Stockholm, Preprint, pp.11--18. Terzaghi, K., 1946. Section I in: R.V. Proctor and T.L. White (Editors), Rock Tunneling with Steel Supports. The Commerical Shearing and Stamping Co., Youngstown, Ohio. Wickham, G.E., Tiedemann, H.R. and Skinner, E.H., 1972. Ground support prediction model (RSR concept). Proc. Rapid Excavation and Tunneling Conf., AIME, Chicago, 1: 43--64.