Applications of Rock Mechanics

Applications of Rock Mechanics
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10

Applications of Rock Mechanics in Engineering Geology LEARNING OBJECTIVES After studying this chapter, the reader will be familiar with the following:

• Importance of rock mechanics in engi• •

neering geology and civil engineering works Laboratory testing of common properties of rocks Instrumental measurement of in-situ stress of rock mass

• Methods to determine shear strength and compressive strength

• Rock mass classification of Norwegian •

Geotechnical Institute and geomechanics classification Solving practical problems of support requirement for underground structures

10.1 INTRODUCTION This chapter deals with the methods of quantitative evaluation of rock properties and the elastic and plastic behaviour of rocks under stress. It explains with illustration the various types of instruments used for determining the strength properties of intact rocks and rock mass. The chapter elucidates different approaches for estimating the rock quality designation. It also explains the rock mass classification according to the Norwegian Geotechnical Institute (NGI), as well as the geomechanics classification giving examples of their uses to identify the support requirement of underground engineering structures. It further describes the method of calculation of geological strength index (GSI) of tectonically disturbed rocks such as the Himalayan terrain using the two tables provided.

10.2 RELEVANCE OF ROCK MECHANICS IN EVALUATING ROCK AND ROCK MASS PROPERTIES The study of the physical characteristics and mechanical behaviour of rocks in response to the forces imposed on them comes under the purview of rock mechanics. Application of the principles of rock mechanics is necessary in engineering geological works related to civil engineering structures, for example, concrete and masonry dams, tunnels, and underground powerhouses that are built in or of rocks.

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‘The engineering geologist must recognize that for his work to be of maximum benefit to the design engineer, it will be necessary for the engineering geologist to quantify his findings. That is, engineering geologic results must not only be presented in a descriptive form that is understandable by the engineer, but the engineer now is requesting numerically defined limits for engineering geologic description’, was the message of Judd (1969) in the inaugural ceremony of the symposium on the role of rock mechanics in engineering geology held in India in 1968. In fact, knowledge of rock mechanics is necessary in engineering geological works to understand the behaviour of rocks under force field and to estimate the rock properties in quantitative terms for engineering design. The science of rock mechanics is based on the engineering principles used in the analysis of rock and rock mass for engineering purposes. Rocks are used for engineering purposes mainly in two ways—one as construction material involving only intact rocks and, two, as foundation to engineering structures on rock mass. Road metals, railway ballasts, concrete aggregates, cut-stones, masonry works, support columns, and beams are only few of many instances of engineering use of varied sizes of intact rocks as construction materials. Construction of heavy structures such as high-rise buildings and concrete dams needs a foundation of firm in-situ rocks. The quantitative values of rock properties derived by conducting various tests on intact rocks and measurement of relevant features in rock mass find importance in the engineering analysis of foundation condition, design of slopes, and underground excavation in rocks. In case the rocks are used as construction materials, intact rock specimens are taken from the field site and tested for determination of rock properties in a laboratory set-up. Several instruments are also available to measure the strength properties of intact rocks in the field itself (Fig. 10.1). However, it is the rock mass properties that are important to engineers and involve the determination of the bulk strength properties for foundation of engineering structures and excavation purposes. The rock mass properties are in a way controlled by planes of discontinuities and weak structural features of rocks including faults, fractures, joints, bedding planes, foliation planes, and clay seam. Field measurements of the geological parameters related to these planes of discontinues and other weak features are the main considerations along with the analytical data of laboratory tests on rock specimens in estimating rock mass properties. Wherever needed for design purposes, rock mass properties such as bearing strength of foundation rock and stress condition of rocks at depths are measured by in-situ instrumental Fig. 10.1 Determining the strength of intact rock by tests. point load testing machine

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This chapter deals with the basic aspects of rock mechanics such as quantitative evaluation of the properties of intact rocks as well as in-situ rocks or rock mass having relevance in the construction of surface and subsurface engineering structures with which engineering geological works are closely associated. Brief descriptions of such instruments including their functions are also provided in this chapter.

10.3 DETERMINATION OF COMMON PROPERTIES OF ROCKS The common rock properties such as specific gravity, porosity, void ratio, and absorption can be determined in a laboratory set-up with an oven, an accurate balance, and some glassware. Brief descriptions of the testing procedures for determining these rock properties in intact specimens are given in the following subsections.

10.3.1 Specific Gravity Specific gravity of a rock specimen is defined as the ratio of the weight of the specimen at a given temperature to the weight of an equal volume of water (that weighs1gm/cm3). The procedure to determine the specific gravity in the laboratory is as follows: The specimen is oven-dried for 24 hours and cooled, and its weight (W0) is taken. It is then soaked in distilled water for 24 hours and its weight (Ww) is noted. Finally, the specimen is immersed in water and its weight (Ws) is taken under suspended condition. The specific gravity (G) of the rock specimen is then given by G=

W0 Ww

Ws

(10.1)

The specific gravity thus obtained is the apparent specific gravity of the rock. In igneous and metamorphic rocks—in which the pore spaces are negligible—the apparent specific gravity is almost the same as the true specific gravity. However, in a porous sedimentary rock, the apparent specific gravity will vary to a certain extent depending upon the volume of pore spaces. The true specific gravity of a rock specimen can be obtained by powdering the specimen and then following the method described under ‘specific gravity’ for soil (see Section 6.3).

10.3.2 Density Density is defined as the mass per unit volume. The density ( r) of a rock specimen is derived by dividing the weight of the specimen by its volume. Density is determined in the same way as specific gravity, that is, by measuring the dry weight (W0), water-saturated weight (Ww), and water-suspended weight (Ws). However, unlike the specific gravity, which is a dimensionless number, density has a unit and can be expressed as follows: r=

W0 gm/cm 2 Ww − Ws

(10.2)

Thus, the density of a material is the same as the specific gravity when expressed in metric units. The specific gravity of granite is 2.68, its density is 2.68 gm/cm3 = 2680 kg/m3 = 2.68 tonnes/m3. Density may be of the following different types:

• Dry density, rd is the weight of dry specimen with pores free of water/unit volume. • Saturated density, rs is the weight of the specimen soaked in water/unit volume.

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• Grain density, rg is the weight of the powdered sample/unit volume. • Bulk density, rb is the weight of the specimen with pores partially filled/unit volume. In a porous sedimentary rock, saturated density varies to a great extent from dry density due to pore spaces. The strength of a rock also reduces when its density is reduced due to the presence of void spaces.

10.3.3 Unit Weight In civil engineering works, it is desirable to use the term ‘unit weight’, which is the same as density when expressed in the unit of metric system. Thus, unit weight of basalt (G = 2.65 or r = 2.65 gm/cm3) is 2.65 gm/cm3 or 2650 kg/m3. In English system of measurement as used in the US, unit weight is expressed as pound (lb) per cubic feet (ft3). The unit weight of a rock sample is derived by multiplying the specific gravity of the sample by the density of water, that is, 62.4lb/ft3. Thus, the unit weight of basalt (G = 2.65) in English system of measurement will be (2.65 × 62.4lb/ft3) = 165 lb/ft3.

10.3.4 Porosity Porosity (h) of a rock specimen is the volume of voids contained in it and is expressed as the percentage of the gross volume (V ) of the specimen. In the determination of porosity, if a rock specimen of regular shape is used, the volume (V ) can be directly measured. A rock cube or a rock core with parallel cutting of two ends is generally used to facilitate direct measurement of volume. The specimen is first oven-dried for 24 hours at a temperature of 105°C and then its weight (W0) is taken. It is then kept immersed in distilled water for 24 hours and the weight (Ww) is noted. Porosity is given by the following equation: h=

Ww − W0 × 100 V

(10.3)

A rock with low porosity has high density and possesses high strength. For example, granite, dolerite, charnockite, gneiss, and massive quartzite that have negligible porosity are dense rocks possessing high strength.

10.3.5 Absorption Absorption is the ratio of the weight of water filling the pores in a rock specimen to the weight of the dry specimen expressed as percentage weight. To measure the absorption of a rock specimen, it is oven-dried for 24 hours and its weight W0 is taken. The specimen is then kept immersed in water for a period of 72 hours and the weight W1 is noted. The absorption of the specimen is estimated by the following expression: W1 W0 × 100 W0

(10.4)

A rock specimen kept immersed in water for sufficient time may not absorb water to fill all the pore spaces of the rock due to air in the pores. Some clay present in the pore spaces may swell when the specimen is immersed in water and hinder entry of water into the pores. Thus, absorption is not essentially dependant on the total volume of pore spaces in a rock specimen but depends on its capacity to absorb water. Rocks with high absorption values are generally low in strength (see the tables given in Section 10.4.6).

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10.4 MEASUREMENT OF STRENGTH OF INTACT ROCK Rock strength is the most important parameter in the design of a structure. The stability of a structure depends upon the strength of the foundation rock and its behaviour under stress. The strength of a rock can be assessed by subjecting it to any of the three stresses, namely compressive stress, shear stress, and tensile stress and studying the resistivity of the rock as follows: Compressive strength Compressive stresses comprising two opposite forces applied on a rock specimen act to decrease the volume of the rock specimen. Compressive strength is the maximum stress that is necessary to break a loaded specimen of rock. It is measured as the total load applied per unit area in kg/cm2. Shear strength Shearing action is caused by two forces acting in opposite directions along a plane of weakness (e.g., fracture, fault, bedding plane) inclined at an angle to the forces. It tends to move one part separated from the other part with respect to each other. Tensile strength When a rock specimen is placed under tensile stress, its volume decreases due to the forces directed outwards, opposite in action. The stresses tend to produce cracks in the rock. Tensile strength is lower than compressive strength. The laboratory measurement of the strength of a rock is taken in intact specimen in the form of a cube, cylinder (generally in rock core), or disc. The laboratory equipped with rock cutting and rock drilling machines facilitates preparation of the specimen for the test. A laboratory type drilling machine can drill cylindrical specimens generally of diameters 25 mm, 38 mm, and 63 mm. The rock cutting machine has an arrangement to cut the two opposite sides of the cylindrical rock specimens perfectly parallel by means of its diamond saw. The instruments used and the methods followed in the determination of strength properties of intact rocks in the laboratory are described in the following subsections.

10.4.1 Rebound Hammer Test Rebound hammer is a handy instrument consisting of a spring-loaded steel hammer (Fig. 10.2). It is used to measure the approximate value of compressive strength of rock in hand specimen and also to assess the uniformity characteristics of rock strength in outcrops. When the plunger housing with the hammer is firmly compressed against a rock surface, the spring-loaded hammer gets released automatically and rebounds after giving an impact of energy, the exact amount can be read out from the scale given in the graph card of the rebound hammer. Usually, an energy impact of 0.075 m-kg is satisfactory. Scale for measuring The rebound of the hammer is related to the rebound distance Plunger ultimate compressive strength and modulus of elasticity of rock and is indicated in a scale as rebound number. The rebound distance of the hammer with respect to the plunger is measured from the scale attached to the frame of the hammer. The ultimate compressive strength Graph-chart for reading of the rock is thus obtained within 75 per cent compressive strength confidence limit. For example, rebound numbers 49, 61, and 65 in the standard model of Schmidt Fig. 10.2 Standard rebound hammer hammer are indicative of compressive strengths

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633 kg/cm2 857 kg/cm2, and 935 kg/cm2, respectively. The digital model of rebound hammer automatically calculates the rebound numbers and the compressive strength.

10.4.2 Point Load Test Point load testing machine, shown in Fig. 10.3(a), is a portable instrument used to measure the strength of rock specimens in the laboratory as well as in the field. The instrument consists of a rigid frame, two point load platens that are conical in shape, a hydraulically activated ram with pressure gauges, and a device to facilitate measurement of the distance between the loading points. The pressure gauges of different ranges are given and the one to be fitted with the machine during testing should be of the type that can record the failure pressure. Several sophisticated instruments with digital device for measuring pressure are also available. The rock specimens tested by point load method fail under tension, developing cracks parallel to the loading direction. The point load provides data as point index that can be converted into compressive strength with reasonable accuracy by applying the empirical formula depending upon the nature of the samples. The advantage of this instrument is that it can be used to test irregular samples in the field in addition to testing core and block specimens in the laboratory. Ideally, the samples should be 50 mm in thickness. In general, point load tests are conducted on the following four possible sample types shown in Fig. 10.3(b):

• • • •

Type 1: Core samples for diametrical test Type 2: Core samples for axial test Type 3: Irregular lump sample Type 4: Block sample Point load index (Is) can be calculated using the following relation: Is =

P kg/cm 2 D2

(10.5)

where P is the failure load and D is the distance between the two platens.

Rigid frame Conical platen Pressure guages

50 mm m Sample type 1

Sample type 2

Sample type 3 (a) Fig. 10.3

Sample type 4 (b)

Point load testing: (a) instrument; and (b) different types of samples

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The point load test involves placing the rock sample tightly fixed between the two conical platens and compressing until failure occurs. In selecting samples for testing of blocks and irregular samples, it is to be seen that the value of D remains at 50 mm or very close to it. Otherwise, a correction is needed. The correction is done by using the size correction chart given with the machine. For computation of uniaxial compressive strength parameters (Ic) from the point load index (Is), the following relation is used: Ic = (14 + 0.175D) × Is

(10.6)

where D is the diameter or distance in millimetres. It has been proved that uniaxial compressive strength calculated from point load index gives reasonably reliable values.

10.4.3 Test for Uniaxial Compressive Strength Uniaxial unconfined compressive strength is the common test conducted in rock specimens in the shape of a cylinder or cube. In general, NX-size (54 mm in diameter) cores obtained from drilling are used for testing. The length of the core specimen is kept at 2 to 2.5 times the diameter by cutting and grinding the two end faces parallel to each other. In its simple form, the uniaxial compressive strength testing machine consists of a hydraulic jack for applying pressure with increasing order. The machine is kept on a solid base and is fitted with pressure gauges to measure the applied pressure (Fig. 10.4). The specimen is placed between two platens (e)

(b) (a) (d) (c) Fig. 10.4 Compressive testing machine: (a) Nx-size rock core; (b) hydraulic jack; (c) solid base; (d) platens; and (e) pressure gauge

and pressure is applied slowly until the specimen crumbles. The crumbling may be in the form of two intersecting cracks at acute angles or a set of parallel to semi-parallel cracks. The applied pressure (P in kg) is noted from the pressure gauge and the compressive strength (s ) expressed as applied pressure per unit area (A cm) is calculated as follows: Uniaxial compressive strength s =

P kg/cm 2 A

(10.7)

10.4.4 Tests for Triaxial Compressive Strength There are several types of triaxial testing instruments. Figure 10.5 illustrates a simple type of triaxial cell in which it is easy to conduct several tests without draining between tests (Hoek 2007). It consists of a steel cylinder with two platens for placing the rock specimen kept in a

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Hardened and ground steel spherical seals Clearance gap for strain gauge wires Mild steel cell body Rock specimen with ground ends and a length-to-diameter ratio of 2 Oil inlet—maximum pressure 700 MPa Strain gauges—if required

Rubber sealing sleeve

Fig. 10.5

Triaxial cell (Hoek 2007)

rubber jacket, and the space of the cylinder surrounding the specimen is filled with oil that acts as lateral pressure. The height of the cylindrical rock (generally NX size rock core) specimen that is to be tested should be 2 to 2.5 times its diameter. Hoek’s cells of other sizes such as AX and BX are also available. The oil that surrounds the rock specimen acts as the confining pressure. Thus, in triaxial test, unlike uniaxial test, the rock specimen is subjected to lateral pressure in addition to vertical (axial) stress. The axial pressure is exerted through the top platen. The axial and confining loads can be increased simultaneously, and then keeping the confining stress constant, the axial stress is increased until the failure of P Platen the specimen occurs. For rock specimens having high moisture content (75 per cent Core speccimen or more), it is necessary to measure the pore pressure. The strain gauge fitted with the load cell measures the ultimate strength (the P Platen ability of rock without yielding to break). This strength of the rock depends upon various factors such as the nature of rock including its composition, grain size, and angularity and is also related to the increased loading and fluid content in rock pores. Using the test results on various axial loads and lateral pressures until failure, Mohr’s diagram and failure envelop for the tested rock specimen are plotted to obtain the values of cohesion c and internal friction angle f as done in triaxial soil tests (see Section 6.12,Fig. 6.20). Fig. 10.6 Instrument (HEICO) for Brazilian test for tensile strength showing core specimen of diameter D and length L (L /D
10.4.5 Brazilian Test for Tensile Strength Brazilian test for tensile strength is conducted by applying diametrical compression to induce tensile stress in a thin disc of rock core (Fig. 10.6). The ratio between the length (L) and diameter

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(D) of the rock core test specimen should be less than one (thus L /D
(b)

s1 =

Fig. 10.7 Brazilian test: (a) applied load on specimen diameter; and (b) tension cracks

2P p DL

(10.8)

where s1 is the Brazilian tensile strength (MPa), P is the load at failure, and D is the diameter (mm). Tensile strength of a rock is lower than its compressive strength. A rock under tension will fail earlier to compression. Hence, in the construction of an engineering structure on rocks, it is important to know the stress regime of the rock, especially whether tension will be the primary force.

10.4.6 Test Results on Engineering Properties of Various Types of Rocks The common engineering properties such as density, absorption, and uniaxial compressive strength of major rock types are presented in Tables 10.1–10.3. The results are based on large numbers of specimens of each rock type that were collected in different times from various project sites and tested in the geotechnical laboratories of Geological Survey of India (Gangopadhyay 1990). It may be seen that, in general, igneous and metamorphic rocks are denser and possess higher strengths compared to the sedimentary rocks, except the Vindhyan sandstone and limestone (Table 10.2). Table 10.1

Engineering properties of igneous rocks

Rock type

Density (kg/cm3)

Absorption* (%)

Compressive strength (kg/cm2)**

Medium-grained granite

2.66–2.72

Negligible

1850–4150

Basalt (massive)

2.69–2.92

0.03–1.19

1410–4550

Basalt (amygdaloidal)

2.47–2.87

0.68–4.49

1135–2109

Dolerite

2.78–3.03

Negligible

2550–5400

Gabbro

2.81–2.95

Negligible

2160–5250

Charnockite

2.68–3.01

Negligible

1870–3850

* 24 hours of saturation

Table 10.2

** Dry and fresh rock

Engineering properties of sedimentary rocks

Density (kg/cm3)

Absorption* (%)

Compressive strength (kg/cm2) **

Sandstone (Vindhyan)

2.54–2.60

0.85–1.94

550–1975

Sandy shale (Vindhyan)

2.56–2.65

0.30–1.34

415–682

Limestone (Vindhyan)

2.75–2.76

0.25–0.26

766–1399

Rock type

(Contd )

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Table 10.2 (Contd)

Density (kg/cm3)

Absorption* (%)

Compressive strength (kg/cm2) **

Sandstone (Gondwana)

2.41–2.59

1.15–5.77

90–525

Sandstone (Lower Siwalik)

2.42–2.68

1.88–2.90

450–1060

Sandstone(Middle Siwalik)

2.40–2.47

2.50–2.93

70–250

Sand rock (Upper Siwalik)

2.31–2.41

2.42–3.75

26–240

Shale/claystone (Siwalik)

2.66–2 .77

7.59–9.38

89–123

Sandstone (Tertiary)

2.37–2.59

2.50–5.53

50–160

Limestone (Tertiary)

2.66–2.76

0.15–3.99

261–479

Limestone (Cuddapah)

2.67–2.70

0.02–0.42

335–1350

Shale (Cuddapah)

2.56–2.70

0.41–2.57

144–342

Sandstone (Cretaceous)

2.10–2.32

6.60–7.10

65–350

2.71

3.9–18.0

24–290

Rock type

Tuff breccia * 24 hours saturation

Table 10.3

** Dry and fresh rock

Engineering properties of metamorphic rocks

Rock type

Density (kg/cm3)

Absorption* (%)

Compressive strength (kg/cm2)**

Granite gneiss (medium grained)

2.64–2.68

Negligible

1715–3580

Hornblende–mica gneiss

2.68–2.70

0.50

346–1338

Quartzite (massive)

2.66–2.67

Negligible

550–2850

Quartzitic phyllite

2.63–2.66

Negligible

505–2250

Shaley phyllite

2.32–2.62

0.46–5.32

193–441

Meta-dolerite

2.78–3.10

0.61–1.42

695–2280

Marble

2.66–2.73

Negligible

440–930

Mica schist

2.59–2.60

3.0

176–626

Khondalite

2.60–2.67

0.72–0.95

456–603

* 24 hours saturation

** Dry and fresh rock

10.5 ELASTIC PROPERTIES OF ROCKS In case of deformation under continuously increasing stress on a rock body, it may be found that at a certain stage, the body returns to its original shape if the stress is removed. In this stage, the body is said to have elasticity and the strain is proportional to stress. In addition, a strained elastic material stores the energy used to deform it, and the energy is recoverable. The body is said to have reached its elastic limit at the stage when the magnitude of strain begins to exceed the magnitude of stress permanently. When the strain goes beyond the elastic limit of the rock, plastic flow takes place. If the rock is constituted of brittle materials, the plastic flow will be small in extent, but in case of the rock that is ductile in nature, the plastic flow will be large. This elastic behaviour of a rock is related to stress (s ) and strain (e). Stress is measured by the relation s = P/A, where P is the force exerted in intact rock in an area A. Strain is given by the expression e = ΔL /L, where L is the length and (ΔL) is the change in length of rock specimen.

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The ratio between stress and strain is known as the modulus of elasticity or Young’s modulus (E) and is expressed as s P A = =E e ΔL L L

(10.9)

If the lateral stress in a rock is given by B, the ratio between strain of a material in lateral extension (lateral strain) to the strain under vertical extension (axial strain) designated as Poisson’s ratio is given by the following relation: Lateral strain ΔB B B = = m (Poisson’s ratio) Axial straim ΔL L L

(10.10)

Poisson’s ratio ( m) in a rock varies between 0.1 and 0.5. During earthquake, the waves move through the rock guided by the elastic properties of the rock. The velocity of wave propagation depends on Poisson’s ratio, which is variable in different rocks. Thus, depending upon the nature of rock, the wave velocity generated by earthquake will change. In the same rock too, the property may vary depending upon the interlocking nature of the mineral grains. Rocks with rigid interlocking grains will have high modulus of elasticity, but with a large content of moisture as in the case of an immersed body of rock, the modulus of elasticity will be reduced. This aspect of elastic behaviour is significant in the design of engineering structure to be founded on rock. Modulus of elasticity or Young’s modulus is a measure of the rock property that resists deformation. When a cylindrical specimen of rock is subjected to stress parallel to its long axis, it will lengthen and the diameter will be under tension. Poisson’s ratio, that is, the ratio of lateral strain to axial stress is measured when a cylindrical rock specimen is subjected to compression parallel to the axis of the rock specimen; the rock shortens along its axis while its diameter increases.

10.6 MEASUREMENT OF STRESS IN UNDERGROUND ROCKS The stability of an underground opening depends on the rock mass strength and the stress that existed in the rock before the excavation. Tectonic events during geological time are responsible for inducing this in-situ stress in rock. The magnitude of this pre-existing stress varies widely depending upon the nature of geological history of the rock formation in which the in-situ stress exists. The measurement of this in-situ stress in underground rocks is necessary before excavation is taken up for constructing an underground structure. The test results are utilized in the design of the underground structures and finding remedial measures against any deformation of strata that may be caused by the release of stress under the new condition. In fact, excavation design of all large underground openings such as tunnels, railways, powerhouses, and storages for oil or nuclear waste disposal requires measurement of the in-situ stress in rock mass. The methods of measuring in-situ stresses are varied but two significant methods include the flat jack test and the borehole deformation over-coring method.

10.6.1 Flat Jack Test (Direct Stress Measurement) Flat jack test is the most convenient method for measuring in-situ stress in rocks of underground openings. The flat jack is a thin envelop-like bladder made of stainless steel with inlet and outlet portals that can be pressurized with hydraulic oil. It measures the pressure required to restore the set of measuring pins fixed in rock wall on either side of a slot. Flat jacks are manufactured in different configurations such as square, rectangular, or with curved edges and their sizes also

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(a)

207

(b)

Fig. 10.8 Flat jack with inlet and outlet portals: (a) rectangular; and (b) curved edged

vary to a great extent, generally between 30 cm × 30 cm and 60 cm × 60 cm, but may even be a square metre depending on the application and slot preparation. Flat jacks may be of different shapes (Fig. 10.8) such as square or rectangular but a curved flat jack is generally used to fit a slot cut by a circular saw. The test is based on the principles of stress release phenomenon and elimination of local stresses followed by controlled stress compensation. The test is conducted in three phases (Fig. 10.9) as follows: (i) A small portion of the wall is selected for the test where the rock is not affected by any joint or facture and is also free from any external effects such as cracks due to blasting. (ii) Two sets of stainless steel pins (reference pins) are then fixed in the wall rock keeping a space between them for inserting the flat jack. After setting the pins and measuring the distance (say, s), a slot is cut by means of a diamond saw so that the pins on either side tend to move their positions due to stress relief phenomenon. The process relieves the rock surface of the stress that originally existed across it. This is measured between the slots by means of the pins fixed prior to cutting. If the distance between the pins is s1, then s1
p

s1
s

(a)

p

(b)

pf

s = s1

(c)

Fig. 10.9 Different phases of conducting flat jack tests; (a) before cutting slot; (b) after cutting slot; and (c) after inserting flat jack and applying pressure

The complete set of equipment necessary to conduct this test is shown in Fig. 10.10. For conducting the flat jack test in underground openings, it is necessary that an exploration adit or pilot tunnel is driven up to the exact position or place of the proposed structural cavern. The test is then taken up in a limited area of the excavated rock wall. The virgin stress remains unaffected even after excavation, but if the wall rock is extremely fractured, either by tectonic deformation or by blasting effect, the test cannot be effectively done.

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10.6.2 Borehole Deformation Over-coring Method of In-situ Stress Measurement

Fig. 10.10 A set of flat jack testing equipment including a flat jack, a hydraulic pump (used for inflating flat jack) with pressure gauge, hose pipe, and guide frame for positioning of reference pins and slot (HEICO)

In-situ stress in rock proposed for construction of an underground structure can be measured as follows: A small borehole is drilled to the desired depth. Then, the instrument containing strain gauges is inserted through the borehole and the initial reading is taken. Then, another large hole is drilled outside the small borehole to relieve the stress around this drilled part and the strain gauge reading is taken again. The difference in readings provides the measure of the borehole deformation. The techniques used are explained with illustrations (Figs 10.11–10.13) in the following steps:

(i) To start with, a large diameter (NX size) diamond drill hole is drilled to the depth zone at which stress measurement is desired. After recovery of the drill cores, the bottom of the hole is flattened by a special drill bit as shown in Figs 10.11(a) and (b). A smaller diameter (EX size) borehole is then drilled further down from the end of the large diameter hole, see Fig. 10.11(c). NX(76mm) EX(36mm) NX (a) (b) (c) Fig. 10.11 Borehole over-coring method: drilling of holes

(ii) The measuring cell containing a number of strain gauges is then inserted with a special installing tool having an oriented device and a cable to read out unit. Compressed air is used to expand the cell in the hole and the strain gauges are cemented to the wall rock of the small borehole as shown in Fig. 10.12(a). The measuring cell now is fixed to the hole and initial reading (0 reading) is taken, see Fig. 10.12(b). Compressed air

Cable (a)

(b) Fig. 10.12 Borehole over-coring method: insertion of measuring cell

(iii) The installing tool is then removed and the small hole is over-cored by a large diameter thin-walled diamond bit, thus relieving the stress at the core. The corresponding strains are

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recorded by the strain gauge rosettes, see Fig. 10.13(a). The core is then recovered with a special core catcher, see Fig.10.13(b), and immediately after removal, the second reading is taken. From the recorded strain, the stress is computed from laboratory determination of the elastic moduli of rock cores (see Section 10.5).

(a)

(b)

Fig. 10.13 Borehole over-coring method: Stress measurement

This method may not be suitable if the boreholes are very deep as it is difficult to handle the instrument at great depths. In boreholes deeper than 50 m, the hydraulic pressure technique is generally followed for estimating in-situ stress in rocks. The principle of this method is to determine the magnitude of the in-situ stress by the fracture pressure and the sealing pressure. The fracture pressure is needed to produce cracks in the borehole walls and the sealing pressures are needed to maintain the cracks when the pump is stopped. In this method, at the required depth, a section of the borehole is sealed by two rubber packers. Hydraulic pressure is applied to the internal walls between the two packers. When the breakdown pressure is reached, the rock surrounding the borehole fails in tension and develops cracks. This fracture is extended away from the boreholes by continuous pumping. When the pumps are shut off with the hydraulic circuit kept closed, a shutdown pressure is recorded. This pressure is necessary to keep the cracks open. The breakdown and shutdown pressures are related to the virgin stress in the site. A borehole camera can be used to measure the direction of the cracks. Thus, both the magnitude and the direction of the principal stress can be estimated.

10.6.3 Borehole Extensometer Test for Measuring Rock Movement The rocks in an excavated underground cavern tend to move towards the centre of the opening caused by induced stress exceeding the uniaxial compressive strength of the surrounding rocks. The rate of such movement of an underground structure can be measured by a simple instrument named borehole extensometer (Fig. 10.14). The instrument consists of a single rod or wire called single position extensometer that extends between the anchor and the reference head. Extensometer with more than one or two rods (up to a maximum of eight) is known as the multiple position (or multipoint) extensometer (Fig. 10.15).

Dial guage

Rock mass

Grout Reference head Fig. 10.14

Borehole extensometer

Extensometer rods Anchors

Fig. 10.15 A sketch of rod type extensometer

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Both wire type and rod type extensometers may work on mechanical or electrical devices. In practice, the instrument is grouted into the borehole keeping the reference head on the rock wall of the excavated opening. Periodic readings are taken by the sensor on all the points in the reference head. The difference between the initial and the final reading indictes the movement of rock mass during the period. The location, orientation, and length of a borehole extensometer depend on the geotechnical features of the project site and the construction method for determining the depth, direction, and amount of anticipated rock movement.

10.7 ESTIMATION OF ROCK MASS PROPERTIES In the laboratory, only a small portion of the volume of rock mass is considered for testing the rock properties. Hence, it is not representative of the rock mass or in-situ rocks where an engineering structure will be built. Excavation of an underground cavity also involves a large volume of rock mass, and test data on intact rock specimens cannot provide necessary information for design purposes. It is, therefore, necessary to find some procedures to estimate the rock mass properties for design of engineering structures, excavation, and other purposes. Several rock mass classifications are now available, which have been developed by several authors to solve the practical problems related to rock mass properties that satisfy the needs of engineering design. The various geological properties have been considered in devising these classifications.

10.7.1 Rock Mass Classifications It is to be remembered that rock mass classification is essential at the early stage of project planning. As such, engineering geological works are aimed at collecting the field data related to rock properties by measurement on in-situ rocks. This data accompanied by the laboratory test data is fed into the classification system for obtaining the quantitative values for the design. The engineering requirement and the engineering geological information pertain to evaluation of the in-situ characteristics of rocks especially for rock deformation and support purposes and other follow-up measures for the design. Information on water condition and in-situ stress are also necessary. In general, the main purpose of the classifications is the safe design of tunnel and underground structures and support system, but they are also applicable for other underground excavations including mining. Several classifications were devised by a number of authors that include Deere et al. (1967), Proctor (1971), Barton, Lien, and Lunde (1974 and 1975) Barton et al. (1980), and Bieniawski (1974, 1975, 1979, and 1989). Earlier to that, Terzaghi (1946) developed a rock mass classification for the purpose of design of tunnel support in which loads are estimated on the basis of descriptive geology.

10.7.2 Rock Mass Classification of Terzaghi The rock mass classification of Terzaghi (1946) is based on the rock condition in the tunnel. It bears good engineering geological information pertaining to rock quality meant for engineering design of tunnel support (Section 16.10, Table 16.2). Tunnel rocks have been grouped under different categories from hard rock to swelling rock giving the weak features as follows: Intact rock It contains neither joints nor hair cracks and it breaks across sound rock. Blasting in rock may cause spalling of rock from roof. Hard intact rock may face popping conditions of sudden violent outbursts.

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Stratified rock It consists of individual stratum with little or no resistance to separation along boundaries of strata. Transverse joints may or may not weaken the strata and spalling condition may be common in this rock. Moderately jointed rock It contains joints and hair cracks but blocks created by joints will not require support. Both spalling and popping conditions may be encountered. Blocky and seamy rock It consists of rock fragments nearly or entirely separated from each other and as such vertical wall may need lateral support. Crushed but chemically intact rock In this type of rock, the fragments are like sand grains. When crushed rock occurs below water table, it will behave as water-bearing sand. Squeezing rock It advances slowly into the tunnel without any perceptible change in volume. Small micaceous materials or clay minerals of low swelling capacity remain in the rock. Swelling rock It advances into the tunnel because of expansion caused by clay minerals such as montmorillonite having high swelling capacity.

10.7.3 Rock Quality Designation Index The rock quality designation (RQD) index developed by Deere (Deere et al. 1967; D.U. Deere and D.W. Deere 1988) is extremely useful in engineering geological description with quantitative estimate of rock cores obtained from subsurface drilling. In fact, most of the authors have used RQD as one of the important geological parameters in their rock mass classifications. RQD is a measure of lengths of core pieces separated due to joints with respect to total length of rock cores. It is the percentage of core recovery obtained in core pieces of 10 cm or more in length out of the total length of the drill hole run in rock. Thus, RQD(%) =

(Total length of cores of 10 cm or more) × 100 Total length th of core run in cm

Total length of rock core run T 300 cm

L = 48 cm L = 0 cm L = 34 cm L = 43 cm L = 28 cm Fig. 10.16 Measurement of rock cores for determination of RQD

Table 10.4

Consider that in a site, rock mass has been drilled for a length of 300 cm in five runs. Rock cores obtained from the five runs are of lengths 28 cm, 43 cm, 34 cm, 0 cm, and 48 cm (Fig. 10.16). Thus, the total length of cores obtained is 153 cm. Therefore, ( + RQD =

+

+ + ) × 100 = 51% 300 This falls in the category of fair according to Deere’s classification (Table 10.4).

Deere’s RQD classification and description

Rock quality designation (%)

Description of rock quality

0–25

Very poor

25–50

Poor

50–75

Fair

75–90

Good

90–100

Excellent

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Since RQD represents rock mass quality, the pieces of rock cores measured for RQD must be representative of the in-situ condition. The orientation of the drill hole with respect to rock attitudes may cause some splitting of cores and affect the value of RQD. Faulty drilling operation may also result in more fragmentation of rock cores. It is therefore necessary that utmost care is taken to obtain core pieces that resulted from geological causes such as joints and not by artificial reasons. The drill hole should be made to obtain NX size cores using double tube core barrel and diamond bits ensuring that fracturing of cores does not take place by drilling process. When drill hole core is not available, RQD can be computed from the measurement of number of joints in one cubic metre of rock in tunnel or drift by applying the following relation: RQD = 115 − 3.3JV where JV is the sum of the number of joints per unit length for all joint sets. This is known as volumetric joint count.

10.8 NGI ROCK MASS CLASSIFICATION TO ESTIMATE TUNNELLING QUALITY INDEX The rock mass classification (Table 10.5) developed by Barton, Lien, and Lunde (1974) of NGI is a widely used classification in engineering geological works for measuring the rock quality of tunnel and other underground excavations in rocks. The NGI classification estimates the tunnelling quality index (Q) from various geological parameters of tunnel rocks that can be used for the design of support systems.

10.8.1 Empirical Equation Used in NGI Classification The empirical equation adopted in NGI classification to derive rock tunnelling quality index Q the numerical value of which varies from 0.001 to 1000 in a logarithmic scale is as follows: Q=

RQD Q Jr Jw × × Jn Ja SRF

(10.11)

where RQD is the rock quality designation, Jn is the joint set number, Jr is the joint roughness number, Ja is the joint alteration number, Jw is the joint water reduction factor, and SRF is the stress reduction factor. The rock tunnelling quality Q is considered to be a function of the following three measures: (i) Block shear RQD/Jn (ii) Inter-block shear strength Jr/Ja (iii) Active stress Jw/SRF Explaining these three quotients, the authors (Barton, Lien, and Lunde 1974) offered the following comments: ‘The first quotient (RQD/Jn), representing the structure of the rock mass, is a crude measure of the block or particle size, with the two extreme values (100/0.5 and 10/20) differing by a factor of 400. If the quotient is interpreted in units of centimetres, the extreme ‘particle sizes’ of 200 to 0.5 cm are seen to be crude but fairly realistic approximations. Probably the largest

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blocks should be several times this size and the smallest fragments less than half the size. (Clay particles are of course excluded.) ‘The second quotient (Jr/Ja) represents the roughness and frictional characteristics of the joint walls or filling materials. This quotient is weighted in favour of rough, unaltered joints in direct contact. It is to be expected that such surfaces will be close to peak strength, that they will dilate strongly when sheared, and they will therefore be especially favourable to tunnel stability. When rock joints have thin clay mineral coatings and fillings, the strength is reduced significantly. Nevertheless, rock wall contact after small shear displacements have occurred may be a very important factor for preserving the excavation from ultimate failure. Where no rock wall contact exists, the conditions are extremely unfavourable to tunnel stability. The ‘friction angles’ (given in Table 10.5) are a little below the residual strength values for most clays, and are possibly down-graded by the fact that these clay bands or fillings may tend to consolidate during shear, at least if normal consolidation or if softening and swelling has occurred. The swelling pressure of montmorillonite may also be a factor here. ‘The third quotient (Jw/SRF) consists of two stress parameters. SRF is a measure of (i) loosening load in the case of an excavation through shear zones and clay bearing rock, (ii) rock stress in competent rock, and (iii) squeezing loads in plastic incompetent rocks. It can be regarded as a total stress parameter. The parameter Jw is a measure of water pressure, which has an adverse effect on the shear strength of joints due to a reduction in effective normal stress. Water may, in addition, cause softening and possible outwash in the case of clay-filled joints. It has proved impossible to combine these two parameters in terms of inter-block effective stress, because paradoxically a high value of effective normal stress may sometimes signify less stable conditions than a low value, despite the higher shear strength. The quotient (Jw/SRF) is a complicated empirical factor describing the ‘active stress.’ (Hoek 2007) The parameters used in the tunnelling quality index of NGI classification are given in Table 10.5. Table 10.5 Parameters used in tunnelling quality index Q of NGI classification

1.

Rock quality designation

RQD

Very poor

0–25

Poor

25–50

Fair

50–75

Good

75–90

Excellent

90–100

2.

Joint set number

Notes 1. Where RQD ≤ 10 (including 0), a nominal value of 10 is used to evaluate Q. 2. RQD intervals of 5, that is, 100, 95, etc., are sufficiently accurate.

Jn

A.

Massive, no, or few joints

0.5–1.0

B.

One joint set

2

C.

One joint set plus random

3

D.

Two joint sets

4

E.

Two joint sets plus random

6

F.

Three joint sets

9

G.

Three joint sets plus random

12

1. For intersections use (3 × Jn) 2. For portals use (2 × Jn)

(Contd )

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Table 10.5

(Contd)

H.

Four or more joint sets, random, heavily jointed, ‘sugar cubes’, etc.

15

J.

Crushed rock and earth like materials

20

3.

Joint roughness number

Jr

(a) Rock wall contact (b) Rock wall contact before 10 cm shear A.

Discontinuous joint

4

B.

Rough or irregular, undulating

3

C.

Smooth, undulating

2

D.

Slickensided, undulating

1.5

E.

Rough or irregular, planar

1.5

F.

Smooth, planar

1.0

G.

Slickensided, planar

0.5

(c) No rock wall contact when sheared H.

Zone containing clay minerals thick enough to prevent rock wall contact

1.0

J.

Sandy, gravely, and crushed zone thick enough to prevent rock wall contact

1.0

4.

Joint alteration number

Ja

1. Add 1.0 if the mean spacing of relevant joint set is > 3 m. 2. Jr = 0.5 can be used for planar, slickensided joints having lineation, provided the lineations are orientated for minimum strength.

fr degrees (approximately)

(a) Rock wall contact A.

Tightly healed, hard, non-softening impermeable filling

0.75

B.

Unaltered joint walls, surface staining

1.0

(25°–35°)

C.

Slightly altered joint walls, non-softening mineral coating, sand particles, clay-free disintegrated rocks, etc.

2.0

(25°–30°)

D.

Silty or sandy clay coatings, small clay fraction (non-softening)

3.0

(20°–25°)

E.

Softening or low friction clay mineral coatings, that is, kaoline, mica, chlorite, talc, gypsum and graphite, and small quantities of swelling clay (discontinuous coatings, 1–2 mm or less in thickness)

4.0

(8°–16°) Note: Values of fr, the residual friction angle, are intended as an approximate guide to the mineralogical properties of the alteration products, if present.

(b) Rock wall contour before10 cm shear F.

Sandy particles, clay-free disintegrated rock, etc.

4.0

(25°–30°)

G.

Strongly over-consolidated, non-softening clay mineral fillings (continuous,
6.0

(16°–24°)

H.

Medium or low over-consolidation, softening clay mineral fillings (continuous,
8.0

(12°–16°) (Contd)

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Table 10.5 (Contd) J.

Swelling clay fillings, that is, montmorillonite (continuous,
8.0–12.0

(6°–12°)

(c) No rock wall contact when sheared K.

Zones or band of disintegrated or

L.

crushed rock and clay (see G, H,

M.

and J for clay conditions)

N.

Zones or band of silty or sandy clay, small clay fraction (non-softening)

O.

Thick, continuous zones or

10–13

(6°–24°)

P.

bands of clay (see G, H, and J for clay conditions)

13–20

(6°–24°)

Joint water reduction factor

Jw

Water pressure (kgf/cm2)

5.

6.0 8.0 8.0–12.0 5.0

A.

Dry excavation or minor inflow,
1.0

B.

Medium inflow or pressure, occasional outwash of joint fillings

0.66

C.

Large inflow or high pressure in competent rock with unfilling joints

0.5

D.

Large inflow or high pressure, considerable outwash of filling

0.33

E.

Exceptionally high inflow or pressure at blasting, decaying with time

0.2–0.1

F.

Exceptionally high flow or pressure continuing without decay

0.1–0.05

6.

A.

B.

C.

Stress reduction factor

(6°–24°)

10) (>10)

SRF

(a) Weakness zones intersection excavation, which may cause loosening of rock mass when tunnel is excavated

1. Reduce these values of SRF by 25–50% if the relevant shear zones only influence but do not intersect the excavation.

Multiple occurrences of weakness zones containing clay or chemically disintegrated rock, very loose surrounding rock (any depth) 10.0

2. For strongly anisotropic virgin stress field (if measured): when 5 ≤ s1/s3 ≤ 10, reduce sc to 0.8sc and st to 0.8st; when s1/s3 > 10, reduce sc and st to 0.6sc and 0.6st, respectively, where sc is the unconfined compressive strength, st is the tensile strength (point load), and s1 and s3 are the major principal stresses.

5.0

3. Few case records are available where depth of crown below surface is less than span width. Suggest SRF increase from 2.5 to 5 for such cases (see H).

Single weakness zone containing clay or chemically disintegrated rock (excavated depth 50 m)

2.5 (Contd)

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Table 10.5

(Contd)

D.

Multiple shear zone in competent rock (clay free) loose surrounding rock (any depth)

7.5

E.

Single shear zone in competent rock (clay free) (excavation depth
5.0

F.

Single shear zone in competent rock (clay free) (excavation depth > 50 m)

2.5

G.

Loose open joint, heavily jointed, or ‘sugar cube’ (any depth)

5.0

(b) Competent rock, stress problem

sc/s1

st/s1

SRF

H.

Low stress, near surface

> 200

>13

2.5

J.

Medium stress

200–10

13.0–0.66

1.0

K.

High stress, very tight structure (usually favourable to stability, may be unfavourable for wall stability)

10–5

0.66–0.33

0.5–2

L.

Mild rock burst (massive rock)

5–2.5

0.33–0.16

5–10

10–20

Heavy rock burst, massive rock

(c) Squeezing rock, plastic flow of incompetent rock under the influence of high rock pressure

SRF

N.

Mild squeezing rock pressure

5–10

O.

Heavy squeezing rock pressure

10–20

M.

(d) Swelling rock, chemical swelling activity depending on water pressure

SRF

P.

Mild swelling rock pressure

5–10

R.

Heavy swelling rock pressure

10–20

When making estimates of the rock mass quality (Q), the following guidelines should be followed in addition to the notes listed in Table 10.5:

• When borehole core is unavailable, RQD can be estimated from the number of joints per unit volume, in which the number of joints per metre for each joint set is added. A simple relationship can be used to convert this number to RQD for the case of clay-free rock masses: RQD = 115 − 3.3Jv (approx.), where Jv is the total number of joints per cubic metre (0 Jv > 4.5). • The parameter Jn representing the number of joint sets will often be affected by foliation, schistosity, and slaty cleavage or bedding. If strongly developed, these parallel ‘joints’ should obviously be counted as a complete joint set. However, if there are few joints visible, or if only occasional breaks in the core are due to these features, then it will be more appropriate to count them as ‘random’ joints when evaluating Jn. • The parameters Jr and Ja (representing shear strength) should be relevant to the weakest significant joint set or clay-filled discontinuity in the given zone. However, if the joint set or discontinuity with the minimum value of Jr/Ja is favourably oriented for stability, then a second less-favourably oriented joint set or discontinuity may sometimes be more significant, and its higher value of Jr/Ja should be used when evaluating Q. The value of Jr/Ja should in fact relate to the surface most likely to allow initiation of failure. • When a rock mass contains clay, the SRF factor appropriate to loosening loads should be evaluated. In such cases, the strength of the intact rock is of little interest. However, when

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217

jointing is minimal and clay is completely absent, the strength of the intact rock may become the weakest link, and the stability will then depend on the ratio of rock stress to rock strength. A strongly anisotropic stress field is unfavourable for stability and is roughly accounted for as in Note 2 in the table for stress reduction factor evaluation. • The compressive and tensile strengths of the intact rock should be evaluated in the saturated condition if this is appropriate to the present and future in-situ conditions. A very conservative estimate of the strength should be made for those rocks that deteriorate when exposed to moist or saturated conditions.

10.8.2 Practical Example of Using Tunnelling Quality Index Q The evaluation of the value of Q with the ultimate aim of estimating the support requirement consulting Table 10.5 may appear to be a complex job, but it is easy to use in practice. The following is an example of its use in North Koel tunnel in Bihar in granite gneiss (Gangopadhyay and Mishra 1991). The values of the various parameters are as follows: Table 10.6

Estimating support requirement using tunnelling quality index Q

Rock quality

Good to excellent

RQD = 90

Joint sets

Two sets

Jn = 4–6 (average 5)

Joint roughness

Rough, undulating

Jr = 3

Joint alteration

Tight and hard

Ja = 0.75

Joint water

Dry, minor flow

Jw = 1.0

Stress reduction

Competent, single shear

SRF = 2.5

Thus, substituting the values of RQD, Jn, Jr, Jw, and SRF in Eq. (10.11), we have Q = (90/5) × (3/0.75) × (1.0/2.5) = 28.

10.9 GEOMECHANICS CLASSIFICATION BASED ON ROCK MASS RATING The rock mass classification called geomechanics classification (Table 10.7) developed by Bieniawski (1974) is based on rock mass ratings (RMR) fixed for the different characteristics of tunnel rocks. In addition, rating adjustment is needed with respect to orientation of tunnel drive with attitudes of strike and dip of joints.

10.9.1 Parameters Used in Rock Mass Ratings with Tables The six parameters used for the geomechanics classification are as follows: (i) (ii) (iii) (iv) (v) (vi)

Strength of rock RQD Spacing of joints/discontinuities Conditions of joints Groundwater condition Strike and dip orientations of joints

Both point load and uniaxial compressive strength of rocks in megapascal units (1 MPa = 10 kg/sq) obtained from laboratory tests are considered for the parameter ‘strength of rock’. Other parameters are determined from the study of the drill cores and rocks in the drift. The equivalent ratings for each of these six parameters can be estimated from Table 10.7. The summation of the equivalent ratings for the six parameters gives the final RMR. From the value

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of RMR, the rock mass class with respect to its quality can be known (as in Table 10.7C). It may be noted that Bieniawski from time to time (1974, 1975, and 1979) refined his original rock mass classification and the one reproduced here (Table 10.7) is from the publication in 1989. Table 10.7A Classification parameters and their ratings

Uniaxial compressive Point load test strength (Mpa) index (Mpa)

Rating

Drill core quality (RQD %)

Rating

Spacing of joints

Rating

> 250

> 10

15

90–100

20

>2m

20

100–250

4–10

12

75–90

17

0.6–2 m

15

50–100

2–4

7

50–75

13

200–600 mm

10

25–50

1–2

4

25–50

8

60–200 mm

8

5–25

2

3

5

1–5

1

0.1

Damp

10

Slightly rough surface, separation
20

10–25 l/min

0.1–0.2

Wet

7

Slickensided surface or gouge
10

25–125 l/min

0.2–0.5

Dripping

4

Soft gouge > 5 mm thick, or joint open > 5 mm continuous

0

> 125 l/min

> 0.5

Flowing

0

Table 10.7B Rating adjustment for joint orientations (see Table 10.6E)

Joint strike/dip orientation

Very favourable

Favourable

Fair

Unfavourable

Very unfavourable

Tunnels ratings

0

−2

−5

−10

−12

Foundation ratings

0

−2

−7

−15

−25

Slopes ratings

0

−5

−25

−50

− 60

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Table 10.7C Rock mass classes determined from total ratings

Rating

100–81

80–61

60–41

40–21

Class

I

II

III

IV

V

Description

Very good rock

Good rock

Fair rock

Poor rock

Very poor rock

Table 10.7D Meaning of rock mass classes

Class

I

II

III

IV

V

Average stand-up time

10 years for 5 m span

6 months for 4 m span

1week for 3 m span

5 hours for 1.5 m span

10 hours for 0.5 m span

Cohesion of rock mass

> 300 kPa

200−300 kPa

150−200 kPa

100−250 kPa

Friction angle of rock mass

> 45°

45°−90°

35°−40°

30°−35°

Table 10.7E

Effect of joint strike and dip orientation in tunnels

Strike perpendicular to tunnel axis

Strike parallel to tunnel axis

Drive with dip 45°– 90°

Drive with dip 20°–45°

Drive against dip 45°–90°

Drive against dip 20°– 45°

Dip 45°– 90°

Dip 20°– 45°

Very favourable

Favourable

Fair

Unfavourable

Very unfavourable

Fair

Table 10.7F

Dip 0°– 20° irrespective of strike Fair

Guidelines for joint classification

Joint length

Rating

Aperture

Rating

Roughness

Rating

Infilling

5 mm

2

Moderately weathered

3

10–20 m

1

1–5 mm

1

Smooth

1

Soft filling
2

Highly weathered

1

> 20 m

0

>5

0

Slickenside

0

Soft filling > 5 mm

0

Decomposed

0

10.9.2 Practical Example of Using Rock Mass Rating The geomechanics classification for the values of RMR was used for North Koel tunnel project of Bihar. The tunnel was driven in gneiss perpendicular to foliation strike and joint dip 45°. Ratings for the parameters can be obtained from Table 10.7A. The rating adjustment value

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for joint orientation is taken from Table 10.7B through Table 10.7E. The result of study for equivalent rating for each parameter and the total RMR thus obtained is as follows: Table 10.8

Application of rock mass rating in North Koel tunnel project

Parameters

Value/description

Rating

Strength (compressive)

60–90 MPa

7

RQD

90%

20

Spacing of joints

0.3–0.8m

20

Condition of joints

Rough surface, separation
20

Groundwater

None

10

Rating adjustment

Very favourable

0

Total RMR (after adjustment) = 77

The RMR value of 77 falls under category II in the geomechanics classification of rock mass quality, which is graded as ‘good rock’ (Table 10.7C). As stated before, the estimation of Q values for the same tunnel rock following the NGI rock mass classification comes as 28, which also falls in the category ‘good rock’. Comparing the two systems of rock mass classifications, Hoek and Brown—Emperical Strength Criteria (1980) has concluded that the Q values of the NGI system bears the following relation with the RMR system: RMR = 9 loge Q + 44

(10.12)

Substituting the value of Q (28) of tunnel rocks estimated for the rocks of North Koel project in Eq. (10.12), we have RMR = 9 loge 28 + 44 = 9 × 3.332 + 44 = 74, which is very close to the measured RMR value of 77, indicating that both the classifications are equally effective in estimating the rock mass quality. In fact, the design of tunnel and the support system made on the basis of the values of both Q and RMR are found to be very useful in safe tunnelling (Section 16.10).

10.10 GEOLOGICAL STRENGTH INDEX FOR BLOCKY AND HETEROGENEOUS ROCK MASS [Note: The Hoek method of rock mass strength determination briefly discussed in this section (along with tables) will be especially required for professionals working in underground projects comprising tectonically disturbed rock formation, such as the Himalayan terrain where the rock formations are affected by several faults and thrust. The related publications of Hoek (reference given in this section) may be studied thoroughly for the application of the method in project works.] In a tectonically disturbed terrain such as Himalayan areas, the rock mass may be very blocky containing discrete blocks interlocked by matrix. The strength of such blocky rock mass depends upon the nature of small blocks, interlocking grains, and also their angularity and roughness. These blocks may also be altered to various extents and the interfaces between the blocks may have slickenside or may be filled by clay. In such highly disturbed rocks, the rock quality cannot be evaluated from the NGI and geomechanics methods. The strength of such in-situ rock mass must be evaluated from geological observations and from the test results on individual rock pieces or rock surfaces that

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221

have been removed from rock mass. This problem has been extensively discussed by Hoek and Brown—Empirical Strength Criteria (1980). To estimate the strength characteristics of such highly disturbed rock mass, later Hoek and Brown (1997) proposed the geological strength index (GSI), which was further refined by Marinos and Hoek (2001) who published two charts for estimating the GSI— one for blocky rock mass and the other for heterogeneous rock mass. These have been reproduced here in Tables 10.9 and 10.10, respectively.

VERY R BLOCKY – interlocked, partially disturbed mass with multi-faceted angular blocks formed by 4 or more joint sets BLOCKY/DISTURBED/SEAMY – folded with angular blocks formed by many intersecting discontinuity sets. Persistence of bedding planes or schistosity DISINTEGRATED A – poorly interlocked, heavily a broken rock mass with mixture of angular and rounded rock pieces LAMINA NATED/SHEARED – Lack of blockiness due to close spacing of weak schistosity or shear planes

DECREASING INTERLOCKING OF ROCK PIECES

BLOCKY – well interlocked undisturbed rock mass consisting of cubical blocks formed by three intersecting discontinuity sets

FAIR F Smooth, moderately weathered and altered surfaces

GOOD Rough, slightly weathered, iron stained surfaces

DECREASING SURFA F CE QUALITY U

STRUCTURE R INTA T CT OR MASSIVE – intact rock specimens or massive in situ rock with few widely spaced discontinuities

VERY R GOOD Very rough, fresh unweathered surfaces V

From the lithology, structure and surface conditions of the discontinuities, estimate the average value of GSI. Do not try to be too precise. Quoting a range from 33 to 37 is more realistic than stating that GSI = 35. Note that the table does not apply to structurally controlled failures. Where weak planar structural planes are present in an unfav f ourable orientation with respect to the excavation face, these will dominate the rock mass behaviour. The shear strength of surfaces in rocks that are prone to deterioration as a result of changes in moisture content will be reduced is water is present. When working with rocks in the fair to very poor categories, a shift to the right may be made for wet conditions. Water W pressure is dealt with by effecti f ve stress analysis.

SURFA F CE CONDITIONS

GEOLOGICAL STRENGTH INDEX FOR JOINTED ROCKS (Hoek and Marinos, 2000)

VERY R POOR Slickensided, highly weathered surfaces with soft clay coatings or fillings

Characterization of blocky rock masses on the basis of interlocking and joint conditions

POOR Slickensided, highly weathered surfaces with compact coatings or fillings or angular fragments

Table 10.9

90

N/A

N/A

80 70 60 50

40

3 30

20

10 N/A

N/A

A. Thick bedded, very blocky sandstone The effect f of pelitic coatings on the bedding planes is minimized by the confinement f of the rock mass. In shallow tunnels or slopes these bedding planes may cause structurally controlled instability. B. Sandstone with thin interlayers of siltsone

C. Sandstone and siltstone in similar amounts

C, D, E and G – may be more or less folded than llustrated but this does not change the strength. Tectonic deformation, faulting and T loss of continuity moves these categories to F and H. G. Undisturbed silty or clayey shale with or without a few very thin sandstone layers

: Means deformation after tectonic disturbance

VERY R POOR – V Very smooth slickensided or highly weathered surfaces with soft clay coatings or fillings f

FAIR – Smooth, moderately F weathered and altered surfaces

GOOD – Rough, slightly weathered surfaces

VERY R GOOD – V Very rough, fresh unweathered surfaces

SURFA F CE CONDITIONS OF DISCONTINUITIES (Predominantly bedding planes)

COMPOSITION AND STRUCTURE R

70

A 60

D. Siltstone or silty shale with sandstone layers

E. W Weak siltstone or clayey shale with sandstone layers

F. T Tectonically deformed intensively folded/faulted, sheared clayey shale or siltstone with broken and deformed sandstone layers forming an almost chaotic structure H. T Tectonically deformed silty or clayey shale forming a chaotic structure with pockets of clay. Thin layers of sandstone are transformed into small rock pieces.

B

C

D

E

40

30

F 20

G

H

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Engineering Geology

From a description of the lithology, structure and surface conditions (particularly of the bedding planes), choose a box in the chart. Locate the position in the box that corresponds to the condition of the discontinuities and estimate the average value of GSI from the contours. Do not attempt to be too precise. Quoting a range from 33 to 37 is more realistic than giving GSI = 35. Note that the Hoek-Brown criterion does not apply to structurally controlled failures. f Where unfavo f urably oriented continuous weak planar discontinuities are present, these will dominate the behaviour a of the rock mass. The strength of some rock masses is reduced by the presence of groundwater and this can be allowed for by a slight shift to the right in the columns for fair f , poor and very poor conditions. W Water pressure does not change the value of GSI and it is dealt with by using effecti f ve stress analysis.

GSI FOR HETEROGENEOUS ROCK MASSES SUCH AS FLYSCH L (Marions. P and Hoek. E, 2000)

POOR – V Very smooth, occasionally slickensided surfaces with compact coatings or fillings f with angular fragments

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Table 10.10 Estimate of GSI for heterogeneous rock masses such as flysch (Marinos and Hoek 2001)

Applications of Rock Mechanics in Engineering Geology

223

The character of rocks, whether blocky or very blocky, seamy, disintegrated, sheared, or of heterogeneous nature is to be decided first for the application of the GSI system. In addition, the following properties of the rock pieces are required for use in the Hoek–Brown criterion with the ultimate purpose of application in design of underground excavation and support system for highly disturbed as well as heterogeneous types of rock mass. Uniaxial compressive strength of intact rock mass Hoek–Brown constant Geological strength index Hoek–Brown constant Hoek–Brown constant Deformation modulus

s ci m1 GSI mb s Em

Here s1 and s3 are the maximum and minimum effective principal stresses at failure, respectively, mb is the Hoek–Brown constant for the rock mass, s and a are constants that depend upon the rock mass characteristics, and s ci is the uniaxial compressive strength of the intact rock. Wherever possible, the values of these constants are to be determined by laboratory analysis of the results of a set of carefully conducted triaxial tests on core samples. In case the desired tests cannot be done, the assumed values as given by Hoek (2007) for the section ‘Rock properties’ can be consulted. The latest version of Hoek–Brown criterion has been published by Hoek, Carranza–Torres, and Corkum (2002). This article together with the computer programme known as RockLab can be downloaded for free from the Internet at www.rocscience.com for use in implementing the criteria. The method of GSI system and the Hoek–Brown criterion have been successfully used in many underground projects in the world, giving reasonable estimates of a wide variety of tectonically disturbed rocks. In India, the Hoek’s method was successfully utilized in the design and support selection for the underground powerhouse of Nathpa Jhakri hydroelectric projects in the Himalayan terrain of Himachal Pradesh.

SUMMARY • Engineering geological works related to quantitative

evaluation of properties of intact rock and rock mass are very important in the design of engineering structures, especially underground structures. The intact rock properties such as density, porosity, absorption, and strength (compressive, shear, and tensile) are determined in the laboratory using normal wares and simple instruments. A point load testing machine can be taken to the field to measure rock strength even in irregular specimens in the field itself. Sophisticated computerized instruments are available for measuring uniaxial and triaxial compressive strengths of intact rocks. Test results conducted on large number of Indian rocks collected from various project sites clearly bring out the fact that igneous and metamorphic rocks are in general dense and possess high strength

and are suitable to use as building materials in engineering constructions. In-situ rock stress, which is an important engineering property, is measured by various methods including the borehole method by strain technique or deformation technique for shallow depths and hydraulic fracture technique for deeper parts. The fracture technique is based on the application of hydraulic pressure for cracking the rocks. Rocks possess the property of elastic deformation, the study of which has significant bearing in the design of foundation on rock. Rock mass behaves differently from that of intact rocks. Estimation of rock mass properties is essential in the design of rock tunnelling and support system. These properties cannot be evaluated by the study of intact rocks alone. However, several methods are now

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Engineering Geology

available to estimate the rock mass quality based on the geological information on rock mass condition and measures of structural attitudes of rock. Two of the widely used rock mass classifications that provide rock mass quality in quantitative terms for design purposes have been discussed here. Of these, the NGI classification is used in the estimation of tunnelling quality index (Q) but it is applicable also in other underground works such as mining and underground powerhouse. The other rock classification is the geomechanics classification that is based on RMR. The two classifications for estimating the rock quality involve measurements of various geological parameters of in-situ rocks, especially the nature of discontinuities and other weak features and

also laboratory test data on rock samples. These measured parameters on rock mass and rock samples when introduced into the given tables provide the values of Q and RMR. Engineering geological works on rocks of several tunnels of India were carried out and the values of Q and RMR were estimated for design of tunnel excavation. It has been observed that both the classifications are very useful in designing costeffective safe tunnelling. In estimating the strength characteristics of highly disturbed rocks that include blocky and heterogeneous rock mass, the Q and RMR values may not be suitable; instead, the GSI proposed by Hoek and Brown may be used by consulting the two charts given by the authors.

EXERCISES Multiple Choice Questions Choose the correct answer from the choices given: 1. Test is performed in the laboratory for compressive strength by applying compression on: (a) a smooth core sample of height that is double of its diameter (b) diametrical (curved) side of core (c) a sample having rough surfaces on both ends of rock core specimen 2. The type of specimen needed to find tensile strength is: (a) rock core with length double its diameter (b) thin disc of rock core (c) rock core with equal length and breadth 3. The compressive strength of granite is: (a) 1500 kg/cm2 (b) 4000 kg/cm2 (c) 500 kg/cm2 4. The compressive strength of dolerite is: (a) 2000 kg/cm2 (b) 4000 kg/cm2 (c) 5500 kg/cm2 5. The compressive strength of charnockite is: (a) 2000 kg/cm2 (b) 5000 kg/cm2 (c) 1500 kg/cm2 6. The compressive strength of Vindhyan sandstone is: (a) 5000 kg/cm2 (b) 1000 kg/cm2 (c) 200 kg/cm2

7. The compressive strength of tertiary sandstone is: (a) 50 kg/cm2 (b) 150 kg/cm2 (c) 2000 kg/cm2 8. The compressive strength of Cretaceous sandstone is: (a) 50 kg/cm2 (b) 250 kg/cm2 (c) 100 kg/cm2 9. The compressive strength of gneiss is: (a) 1000 kg/cm2 (b) 3000 kg/cm2 (c) 6000 kg/cm2 10. The compressive strength of marble is: (a) 200 kg/cm2 (b) 800 kg/cm2 (c) 1000 kg/cm2 11. The compressive strength of mica schist is: (a) 200 kg/cm2 (b) 800 kg/cm2 (c) 1000 kg/cm2 State whether the following statements are true or false. Mark (a) if true and (b) if false: 12. Poisson ratio is the ratio of the lateral strain to axial strain. 13. In-situ stress in underground rocks can be measured by using flat jack. 14. Stress in subsurface rock can be measured through a borehole by using a borehole deformation gauge. 15. The movement of rocks in an underground chamber can be measured by a borehole extensometer.

Applications of Rock Mechanics in Engineering Geology 16. In the field, compressive strength of rock samples can be measured by point load tester.

Review Questions 1. What is rock mechanics? Discuss the scope of application of rock mechanics in engineering geology. 2. Define and write the equations for the following properties of rocks: specific gravity, density, unit weight, porosity, and absorption. 3. Explain the method of determining the compressive strength of an intact rock specimen using rebound hammer and point load testing machine.

225

4. Write a short account on testing of intact rock specimens for the determination of uniaxial compressive strength. 5. What are elastic and plastic deformations of rock? Define Young’s modulus and Poisson’s ratio and give their equations. How will a rock core specimen act under compression parallel to the axis of the specimen? 6. Write a short account on flat jack test for in-situ stress measurement in massive rock. 7. What do you understand by rock mass properties? What are the purposes of rock mass classification? 8. Name the different groups of tunnel rocks under the rock mass classification of Terzaghi.

NUMERICAL EXERCISES 1.

2.

What is RQD? Explain how RQD is calculated from the study of rock cores obtained from rotary drilling. From a 30 m-deep drill hole, the core recovery for each 3 m run of drilling from the top towards the bottom is as follows: 0.2 m, 0.5 m, 1.5 m, 1.4 m, 1.9 m, 2.5 m, 0 m, 2.2 m, 2.6 m, and 2.9 m. Calculate the rock quality of drill cores (use Table 10.4). State the method of measuring RQD of tunnel rock by volumetric joint count. [Answer: 52.3%, fair] Write the equation for NGI rock mass classification (Q values) used in the measurement of rocks in a tunnel or underground cavity. Calculate the value of Q for the tunnel rocks. [Hint: Use the formula given in Eq. (10.6) for volumetric joint measurement and calculate RQD. Refer Table 10.5 for different joint parameters and stress reduction factors. Suppose the following are the parameters obtained after measurement in the tunnel rocks as hinted: RQD measured from volumetric joint count

115 − 3.3Jv = 115 − 3.3 × (Jv = 12) = 75.4

Joint set number: three joint sets

Jn = 9

Joint roughness number: smooth, planar

Jr = 1.0

Joint alteration number: silty coating

Ja = 3.0

Joint water reduction factor: minor inflow

Jw = 1.0

Stress reduction factor: Single shear zone with clay SRF = 2.5 Now use Eq. (10.11) (Section 10.8) to calculate the value of Q.] [Answer: 12.1] 3.

What is geomechanical classification? Name the person who developed this classification. What does RMR stand for? What are the parameters used to find the value of RMR? Calculate the value of RMR from the following parameters (use Table 10.7): Compressive strength

50 MPa

RQD

76%

Spacing of joint

0.7–1.0

Condition of joints

(slightly rough, separation
Groundwater

Rating adjustment

favourable [Answer: 72]

Answers to Multiple Choice Questions 1. (a) 10. (b)

2. (b) 11. (a)

3. (b) 12. (a)

4. (b) 13. (a)

5. (a) 14. (a)

6. (b) 15. (a)

7. (a) 16. (a)

8.(b)

9. (b)