Anchoring In Rock And Soil.pdf

7

f\

1

w 8 IM

1 <

ja» ZP

iO &0 — anchoring depth[m]

8.0

ΓΝ4

>o NO <\f csT

| c>

3D0

f

3

■«■*.

225

«0

I

.Co

X*A

Λί

S*

TΊ At T T #

0,75

\

Λ< y

p

-c

NO

10

4

r

y

* +i

' \* "T

J? ΤΓΠ

L·*
15 20 25 »~ anchoring depth [m]

Fig. 10-20. Diagram for determining approximately the embedding depth for a line of anchors in dry gravel (φ = 36°)

extraction of anchors inclined from the vertical at angles of 0 to 45°. The anchors were provided with an anchoring plate 80 cm in d ameter, and were placed 1.0 m below the surface of a compacted layer of sand of the following properties: y = 1.65 t/m3, ω = 6.5 per cent., φ = 32°, c = 5 kPa. The resultant values, Pmax (each the arithmetic mean of three identical tests), which caused the anchor to be uprooted, are set out in the graph in Fig. 10-21. For comparison the values calculated according to formula (10-7) are also shown; this assumes that mk = 1, σΓ = — (F being the area of the abutted F

80

base). Moreover, the additional depth of 1 m which appears in the computed embedding depths of anchors is not considered. The increased fixing strength of inclined anchors embedded at the same depth below the ground surface, may be explained by the greater overall surface area at which the shear strength of the soil mass expelled by the anchor base is active. The deformation (lifting) of the sand surface was not found to follow the tendon noticeably when the latter was inclined, buf appeared in the majority of cases to be approximately above the centre of the inclined anchoring plate (Fig. 10-22). It is evident, therefore, that lifting was a consequence of transverse stresses resulting from the compression of the sand by the front of the anchor plate. There is no proportional increase in resistance to anchor extraction with increasing tendon length, or with increasing angles of inclination over 45° from the vertical. From 45 to 90°, the resistance to extraction more and more depends on the resistance of the anchor to lateral deflection, or the resistance of a group of anchors to movement towards the surface. Because of this,

1i

or anct Wem 0 cm

Id* 8

soil

y-rt o f = 32

-

* 1

1 O) C;

«I.I tTO

r

W-10°

0°

I-1 II Co X

20°

30°

Wott5°

deflection of tendon from the perpendicular (deflection of anchoring plqte from the horizontal)

Fig. 10-21. Graph showing the dependence of anchor load on the limit of overall failure of the soil, and the inclination of the anchor (according to A. S. Kananyan)

3 mm

2 1

^^fe--r-^a^ \ A„

shearing surface

\

surface v

Fig. 10-22. Diagram of surface deformation arising from the tensile force of an anchor inclined at 30° from the vertical (according to A. S. Kananyan)

81

the forces introduced into inclined or horizontal anchors must be limited. The length of the tendon, as computed from formula (10-7) and (10-8), should not exceed 1.4 times the embedding depth of the root regardless of the tendon direction. 10.3.2

Anchoring depth in saturated loose soils

When an anchor root is extracted from a saturated loose soil, the volume of the soil does not increase, as is found to occur in unconsolidated soil. Consequently, the radial stress does not increase as there is no wedge-like effect of the pulled-out soil; it is the radial stress that produces any increases in shear strength, as in the case of dry soils. The stress in an otherwise unloaded soil results solely from the soil dead weight minus the effect of uplift; there is a linear increase in this stress with depth. The stress in a vertical direction, at depth h, is given by: *„ = ( ? - 1) h> and that in the horizontal direction by:

where 1 — v (v is the Poisson's number for the soil). Assuming that for all anchor positions the stress acts radially over the entire root circumference, the shear strength of saturated non-cohesive soils in the vertical direction is given by: τν = σΗ. tg φ,

and in the horizontal direction by: τ„ = σν . tg φ. Along an anchor forming an angle ψ with the horizontal, the shear strength is: τ2 = σν. cos φ tg φ + συ sin ψ = σν cos ^(tg φ + tg ψ). In calculating the soil's resistance to root displacement (uprooting of the anchor), it is assumed that the shear strength is active at the lateral surface of a cylinder coaxial with the anchor. The cylinder is of diameter, d, equal to the maximum root diameter, and of length, h, equal to the distance between the root centre and the ground surface; h can be expressed by the relation:

82

p Λ max π . d. x

ι

By substituting for τ in this equation, we obtain, for the required force P and safety factor mk, the necessary embedding depth for a vertical anchor: m / V^ f , M V π . d(y - 1) fc0 . tg φ the depth for a horizontal anchor: h

v=

m

ft =

*'^

,

(10-9)

(10-10)

and the depth for an inclined anchor: h

m

s

n.d(y-

fc^ l)fc„.cos^(tgp + t g ^ ) '

Π0 1 0 " ;

where Α„ is the depth of the root centre below the horizontal ground surface.

10.4 ANCHORING DEPTH IN COHESIVE SOILS

Cohesive soils are less capable, compared with non-cohesive soils, of resisting the uprooting of anchors fixed in them. This fact relates to the specific characteristics of cohesive soils, particularly the presence of bound water, a larger void ratio, and a different mineralogical structure; in any case, it leads the anchor designers to specify longer anchor roots, and consequently longer anchors overall. A longer anchor, however, means more work and higher cost. Experience tells us that improvements in the functional reliability of anchors are not always proportional to the increased cost involved. Moreover, the necessity of extending the anchor root disproportionately leads to reduction or omission of the anchor tendon, so that the anchor looses its original form and takes over the qualitatively inferior function of a tensioned pile. L. Hobst, giving due regard to these facts, conducted a number of experiments in 1958 and 1959 in which anchors were fixed in cohesive soil by means of expanded bulb-shaped roots. The tests were carried out in beds of Brno Pleistocene (Fig. 10-23) loess with the following characteristics: γ = 1,935 kg/m 3 , ω = 26 %, c = 0.5, granulometric curve as shown in Fig. 10-24. The stressing equipment was mounted on bridge beams (see Fig. 10-5) in order to avoid the transfer of the reaction forces on to the soil medium in the immediate vicinity of the anchors, with consequent invalidation of the test results. The anchoring bulbs were created by concrete fillings at the ends of the boreholes, the latter being widened with expanded swinging knives fixed to the end of the drilling tool, or with

83

72 11

Fig. 10-23. Relationship between load-bearing capacity of anchors fixed in loess with 35 cm dia bulb, and fixing depth

72 o

l·

10 9 8

h

8s

11 <

hk

o

3 2 7 0

100

I

sand

100

i

I

ZOO 300 400 500 capatity of anchor

medium I

|~~

fine

du

J I vent\

°/\90 30 70

\

BO

40

ayey irticle»

III

v-llt Nflt

lfr γ| |

30

20

10 0

I

900

■φ

UN

en

DU

L

600 700 800 [kfi]

1

Q5

0.25

0.1 0.05 mm

Fig. 10-24. Granulometric curve of Brno Pleistecene loess in the 0.01 0.005 0.002 locality of anchorage tests

explosive charges (see Fig. 13-67, 13-68). In the course of the tests observations were made on the deformations occurring in the anchors, and the deformations and cracks appearing on the surface of the soil medium in relation to the magnitude of the tensile force. It has been demonstrated that during the loading of a cohesive soil by an expanded anchor root, friction plays an important part as a result of the transverse stresses arising in the compressed soil medium above the root front. Thus, when the required length of the anchor tendon is calculated, it may be assumed that the soil

84

mass affected by the prestressed anchor is cone-shaped, with an apex angle equal to twice the angle of internal friction φ. Friction and internal cohesion act along the flanks of this cone, and the magnitudes of these components vary according to the characteristics of the soil in which the anchors are fixed. At present, the shear strength of the soil at any particular root fixing depth is determined, as stated earlier, from Coulomb's equation by substituting the appropriate values of φ and c: τ = σ tg φ' + c'. Starting from this assumption, a formula has been derived for the calculation of the tendon length, taking into account the effect of increasing friction from the ground surface to the root front, as occurs in the case of dry noncohesive soils (see Section 10.3.1). If cohesion is considered to be uniform over the entire shearing area, then for an individual anchor: / mk3P cos φ _ /1Λ / ΤΤΤ^, Τ^ Γ· (1(Μ2) V π · f(3c + σ Γ . / . cos φ) The soil characteristics are taken as being within the following limits: K=

c = 1 to lOkPa, φ = 10 to 2 5 ° , / = tgcp, Gr = <*kr-7^Z

=

2 0 0 t0

5 0 0

% k P a

'

The ultimate stress, akr, of the soil is obtained from the results of tests, and derived approximately from standard values. When this stress is exceeded at the interface between the soil and the root front, the root starts cutting through the soil. In view of this, the maximum loading, Pkr, of an anchor is given by the expression PKr = akr. F, in which F is the area over which the upward pressure of the expanded root is exerted. The load-bearing capacity of an anchor is directly proportional to the difference between the cross-sectional area of the anchor root and that of the anchor base proper. In less strong soils showing low values of akr, the expanded root base must be designed with the largest possible diameter. Computed values for the fixing depth limits for variously loaded anchors are shown graphically in Fig. 10-25. When the curves obtained in this way are compared with a curve showing measured soil resistance to anchor extraction at various depths, it is evident that the formula for calculating anchoring depth is applicable in practice (see Fig. 10-23). When a series of cables are anchored, the volume of the affected soil contributing to the resistance to anchor extraction decreases with the axial interval, /, between anchors. The necessary anchoring depth h'z is obtained from the equation:

S5

κ=

mkP . cos φ l(2c + f. σΓ) '

(10-13)

Depth values, A', in relation to the axial interval, /, between anchors, and the anchoring force, P, are shown graphically in Fig. 10-26. The formula is valid up to the limit defined by / = h . tg φ (denoted in the graph by a thick broken line); for larger values of /, however, formula (10-12) is

100

200 300

400

500 600

700 600 900 1000 — tension [kN]

Fig. 10-25. Diagram for determining approximately the embedding depth of solitary anchors in cohesive soils, according to formula (10—12). 6^300kPal

f=25°,

C>10kPa

mk*M

100 200 300 WO 500 600 700 800 900 1000 1100 12001300

*~P [kN]

Fig. 10-26. Diagram for determining approximately the embedding depth of a line of anchors in cohesive soils, according to formula (10-13). In the area below the thick broken line the embedding depth is determined as for solitary anchors

86

used. The effect of the pressure of the root front on the soil above the root is to overcome molecular cohesion among the soil grains, extrude water from the pores, and modify the internal structure of the soil by forcing the grains of the compressed soil up against one another. These changes result in major or minor permanent deformation displacement of the root apparent in every stress-strain diagram for an anchor fixed in cohesive soil (see Fig. 10-27). The development of this deformation is a long-term process, which means that the anchors have to be designed and the anchor heads mounted on the anchored structure in such a way as to make possible additional stressing of the anchor up to the time of complete consolidation of the loaded soil (see Chapter 19, Fig. 19-5). The validity of formulae (10-12) and (10-13) depends on the magnitude of the soil shear strength around the root, and this strength is decisively affected by the soil moisture content. In naturally moist loess loams, the formation of a conical shear surface has been observed in the laboratory [86]. Following saturation of the soil (28 per cent moisture) the cone did not form and a local failure occurred in the vicinity of the anchoring plate. The soil was displaced to the underside of the plate and laterally as the plate was pulled out under constant tension, but it was only when the plate was nearing the surface that the conical soil body and the anchoring plate were torn out completely. It is clear that the soil undergoes plastic deformation under the pressure of the expanded root, and recedes to the sides. Initially, the resistance to anchor movement increases sharply until local soil failure occurs; then the resistance remains constant, since by forcing the root through the soil further failures occur continuously, with the soil flowing around the root. L. N. Dzhioyev [46] has observed similar phenomena during field tests on vertical anchors with expanded roots (concrete bulbs), in a soil with the following parameters: y = 17kN/m 3 , ω = 23.8 per cent., tg φ = 0.42, c = 17.7 kPa. Bulbs with diameters of 20, 35 and 50 cm were tested, anchored at depths of 110 and 200 cm. The load characteristics showed three successive stages of deformation (Fig. 10-27); the first part of the curve (up to the deformation of soil

.§ [.depressiony

circum flowing

yextractiony

^, partial extraction of anchor

/f

Fig. 10-27. Load diagram obtained during tensile testing of an anchor with an expanded root (bulb) in cohesive soil (according to Dzhioyev)

87

Fig. 10-28. Results of comparative tests of the load-bearing capacities of anchors in sandy clay 1 — unexpanded borehole, anchor: bar of 32 mm dia; root fixed with poured-in cement slurry, 2 — unexpanded borehole, anchor: pipe 5/25 mm; root grouted under pressure and regrouted, 3 — borehole mechanically expanded, anchor: bar of 32 mm dia with base, root grouted under pressure, 4 — borehole expanded by blast, anchor: bar 32 mm dia, root grouted under pressure, 5 — borehole filled with cement mix and expanded by blast, anchor: bundle of 12 x P 7 wires, 6 — borehole filled with cement mix and expanded by blast, anchor: bundle of two 6-strand ropes 24.5 mm dia splayed into strands in the root section

88

P c r i t point) was linear, corresponding to the phase of increasing soil compression above the anchoring bulb; no failure occurred. The second part represented continuous local soil failure, as the bulb was extracted under a constant force, PCTlt, without any apparent disturbance of the ground surface; in the tests this stage was abbreviated as the ratio of bulb depth to bulb width (hjd) decreased. The third stage commenced when the force necessary to lift out the soil above the root coincided with P c r i t , at which point the ground surface started to deform and the tensile force dropped. P c r i t could be computed from the equilibrium of forces occurring at this point. Tests performed at various sites have shown that the load-bearing capacity of anchors fixed in cohesive soils depends on the width of expansion of the anchor root. It has been confirmed that the fixing of anchors in unexpanded boreholes is of little effect in cohesive soils. An example may be cited of measurements of the load-carrying capacity of horizontal anchors conducted by Geoindustria (Czechoslovakia). The tests were carried out on anchors with expanded and unexpanded roots of identical length (3 m) embedded in sandy clay at a depth of about 2.5 m below the ground surface. Two anchors were inserted in unexpanded boreholes, a third into a borehole expanded with a Böhler — Klemm mechanical reamer, and a fourth into a borehole expanded by letting off two charges in the hole, each consisting of 300 g of Semtex. Two further boreholes were expanded by means of two charges (250 and 150g) which were inserted in the grout together with the anchors and blasted. When the grout in the boreholes had hardened, tensile tests were carried out on the anchors up to their ultimate loads. All the anchor roots were then dug out and investigated. The results of the tests are shown in Fig. 10-28. The greatest load-bearing capacity was attained with roots expanded by blasting in the grout, even where the second cavity was incompletely filled owing to insufficient grouting following blasting. The least load-carrying capacity was that of the anchor which was fixed in the unexpanded borehole with grout poured into it.

Chapter 11 MATERIALS U S E D IN THE CONSTRUCTION OF A N C H O R S ( T E N D O N M A T E R I A L )

Anchor tendons can consist of bars, and exceptionally, of steel pipes; in most cases, however, they are composed of straight patented wires or strands. The use of strands is becoming increasingly widespread in spite of their greater cost over other tendon components. The selection of a suitable type of anchor for securing a particular structure to ground, depends on the required load-bearing capacity, the length and number of anchors needed, and the available facilities for placing, fixing and stressing the anchors at the site. As far as purpose is concerned, the use of the various types of anchorage may be outlined as follows: short bar anchors (bolts) are useful where small tensile forces (up to 100 kN), are to be distributed among a large number of short anchors, for example in the stabilization - of a rock face. Lcv^ ! r anchors (up to 15 m) can be rapidly installed for taking up larger forces (up to 400 kN) where sufficient boring capacity and space for manipulation are available, as for example in foundation pits. Cable anchors, on the other hand, are useful for the transfer of considerable tensile forces from structures to the deeper zones of the bedrock. With regard to manipulation, short anchoring ^bars are the simplest in terms of preparation, placing and prestressing. Longer bars are rather more difficult to handle, and for long anchors, tendons of wire or strands are preferable; the latter are flexible, and are thus more easily transported and inserted into boreholes of several tens of metres long, evenjfroni relatively smalLmanipulating platforms and regardless of borehole direction. Functionally prestressed high quality steel anchor tendons are the most suitable for anchoring structures into rock and soil. They reduce expenditure on steel and minimize the boring and prestressing requirements. Moreover, reductions in prestressing caused by rock creep, and more especially by soil creep, will be diminished. The prestressing of high quality steel up to the yield point, produces an elongation several times greater than that created by a similar stress in a steel bar. Assuming the same modulus of elasticity for all types of steel (E = = 190,000 MPa), the following elongations are obtained by prestressing to the yield point. For bars of steel:

91

«•-■τ-Ä-■"»·"■ For steel wire:

sb = E*L = _J2°_ = 0.0042. *

E

190,000

For steel strands: £s

_ <x0.2 _ 1,440 £ 190,000 "

υ υυ/0

·

·

When a tendon of prestressed strands is used, there is less danger of the prestressing disappearing as a result of ground creep. Seven times more ground creep is allowable (before prestressing in the tendon is lost), where strands are used and prestressed to the yield point instead of steel bars. Successful anchoring of structures and stabilization of slopes depends, therefore, on the use of tendons made of prestressing wire and strands, or prestressing bars of high yield point. Bars of low quality steel are normally suitable only for short anchors with a short service life, such as those used in securing rock surfaces in small underground excavations, or in situations where no prestressing is employed. The required characteristics of the material for prestressed anchor tendons may be divided into principal and complementary characteristics. The principal characteristics are strength (Rm), ductility (A), and relaxation value (Rr). The initial prestressing must be maintained over a long time, and therefore high strength coupled with low relaxation losses are desirable. The capacity to absorb energy is represented by the ductility, which is directly proportional to the safety of the tendon against sudden failure. Among the complementary characteristics the following may be mentioned: resistance to corrosion, absence of twists or kinks, flexibility and contractibility (Z), stress-strain characteristics and modulus of elasticity (E), yield point (Rp 0.2 and/or Rt 1.0), elastic limit (Rp 0.1), and fatigue load-bearing capacity. Some complementary characteristics are closely connected with the principal ones. Thus, for example, the requirements that Rp 0.2/Rm = 1.0, that a long linear section is shown extending from the origin on the stressstrain diagram, and that the value of the ratio Rp 0.01 IRm is high, are related to the requirement for low Rr and high Rm values. Contractibility characterizes the plastic properties of the steel better than the value of ductility (4).

92 11.1 MATERIALS FOR THE CONSTRUCTION OF BAR ANCHORS

Neither plain nor ribbed bars exhibit mechanical characteristics as favourable as those of wires; bars are less flexible, although they do ensure a simple, slip-free threaded anchorage, and show greater resistance to corrosion on account of their smaller surface-to-mass ratio. Until recently bar anchor tendons were made of reinforcement steel, with a yield strength of up to 400 MPa, and a strength limit of 600 MPa for anchors secured with cement; some were made of structural steel, with a yield strength of from 285 to 383 MPa, and a strength limit of from 600 to 700 MPa for anchors fixed mechanically. Recently other steels, with a yield strength of 1,100 MPa and a strength limit of 1,300 MPa have been used. Table 11-1 compares the properties of various steels. The use of steels of low yield point is permissible for short anchors (bolts), which can be fixed simply by forcing a wedge into the split end of the bar (see Fig. 13-4). The fixing of such bars in boreholes with grout or cement is TABLE 11-1 Properties of prestressing and threaded steel bar Description, {Standard)

Min. breaking strength [N/mm2]

Macalloy (England) Macalloy Dywidag (GFR)

1,000

Dywidag Krupp HWR (GFR) JISG 3109-1871 (Japan) Sumitomo (Japan) Neturen (Japan) *) Threaded bar

T1)

Min. yield strength [N/mm2]

1,030 1,030

824—980 835

0 16 mm T1)

1,230 1,470 1,080 1,290

1,080 1,325 930 1,060

A2 A2 Bl B2 T1)

940 1,030 1,080 1,175 1,175

785 785 930 930 930

, 1,225

1,080

Max. relaxation [per cent]

1.5 1.5 1.5 1.5 1.5

93

favourably affected by their larger surface area per unit prestressing force. Furthermore they are cheaper than bolts made of steel of high yield point. For anchors longer than 2 m, and particularly for those anchors intended to maintain permanent anchoring forces or permanent prestressing of rock, bars of reinforcing steel are unsuitable. In such cases prestressing bars with a yield strength of 1,000 MPa or more must be used (Fig. 11-1). Bars used for anchoring have pressed ribs on the surface (Fig. 11-2) and are supplied in 12 m lengths; for special requirements they can be supplied in 18 m lengths, and with diameters of up to 50 mm. In Japan high parameter Neturen bars are in use. In Europe, the Dywidag bars are widely used for anchorages in rock; the currently available diameters are 12.2, 26, 32 and 36 mm, and their strength is somewhat higher than that of the bars listed in Table 11-11. Plain bars are used, and bars with rolled-on threads along their entire length (Fig. 11-3) are increasingly coming into use. The latter are supplied in two qualities, in diameters from 15 to 36 mm (Table ll-III). Bars provided with threads in the rolling mill have some advantages, and are therefore preferred in practice over plain bars; one important factor is that

Fig. 11-1. Macalloy bars with cold rolled threads

Fig. 11-2. Bars with pressed ribs

94

the threading strengthens the bond with the grout in the borehole. In addition, the setting up of anchors based on these bars is easier than with the use of plain bars (see Chapter 12). TABLE 11-11 Technical data for prestressmg bars Bar dia in mms

Sectional area in sq. mms

Breaking load (kN) Macalloy

Dywidag

Sumitomo

9.2

66.2

68.3 (3) 78.1 (4)

11.0

95.1

98.4 (3) 112.4(4)

13.0

132.7

137.3 (3) 156.4 (4)

15.0

176.7

16.0

201.1

17.0

227.0

20.0

314.2

23.0

415.5

25.0

500.0

191

234.5 (3) 267.9 (4)

476.4 (4) 490.9

26.0

530.9 551.5

32.0

804.2

800

35.0

862.1

950

*) ) 3 ) 4 )

534.1 (3) 610.5 (4) 568 (1) 678 (2)

36.0

2

416.7 (3)

325

26.5

40.0

191 296

828 (1) 989 (2)

812.7 (3) 928.8 (4)

1,049 (1) 1,252(2) 1,256.6

1,250

Grade St 835/1030 Grade St 1080/1230 Class A2 Class B2 Fig. 11-3. Prestressing bars with pressed threads along the entire length

95 TABLE ll-III Technical data for Dywidag threaded prestressing bars Bar diameter

mm

Sectional area

15.0

16.0

26.5

32.0

36.0

mm2 176.7

201.1

551.5

804.0

1,017.8

Min. yield strength

MPa 885

1,325

835

1,080

835

1,080

835

1,080

Min. tensile strength

MPa 1,080

1,470

1,030

1,230

1,030

1,230

1,030

1,230

Min. breaking load

kN

296

568

678

828

989

1,049

1,252

191

11.2 PRESTRESSING WIRE

Wire for the preparation of anchor tendons is made of cold-drawn plain carbon steel melted in the Siemens — Martin furnace. The basic heat treatment — the patenting — which gives the material its special structural properties and which provides the necessary strengthening during the cold-drawing, takes place in furnaces with automatic thermal control (i.e. control of the heating temperature and the temperature of the lead bath which constitutes the cooling medium). Heat treatment is followed by the pickling process, in which the material is de-scaled with chemicals and its surface is prepared for the subsequent shaping by cold-drawing. In cold-drawing the wire cross-section is gradually reduced as the wire is extended. At the same time, the tensile strength, the elastic limit and the yield strength are increased. Nowadays an additional heat treatment, referred to as low temperature tempering, is carried out on patented wires and cables made from this material. This treatment at 400 °C produces a further increase in the yield strength, which rises to more than 80 per cent, of the overall strength. In some countries, stabilized prestressing wire is also made. Cold-drawn wires are prestressed at a temperature such that a permanent elongation of the wire takes place. The relaxation losses of these wires are less than 2 per cent. (Fig. 11-4). It should be realized that prestress losses which can appear due to the relaxation of the steel are very small compared with those arising from creep of the rock. It is not necessary therefore, to use stabilized wire for anchors, since any losses of prestress by relaxation are made good in most cases by additional stressing (see Chapter 17). Steel works generally supply the following types of prestressing steels:

96

— Patented, untempered, high-strength smooth steel wire, diameter 2.0 to 7.0 mm, coil weight approximately 120 kg, internal coil diameter 500 to 600 mm (Table 11-IV). TABLE 11-IV Smooth patented wires Standard

Diameter (mm)

Type

tfpt

ASTM A 421-74 (USA)

4.88 4.98 6.35) 7.01 4.98 6.35

WA WA WA WA BA BA

1,725 1,725 1,655 1,622 1,655 1,655

BS 2691-1969 (Great Britain)

4 5 5 7 7 4 5 5 7 7

NR NR NR NR NR LR LR LR LR LR

1,720- -1,950 1,720- -1,950 1,570- -1,800 1,570- -1,800 1,470- -1,700 1,720- -1,950 1,720—1,950 1,570--1,800 1,570—1,800 1,470- -1,700

1,460 1,460 1,330 1,330 1,250 1,550 1,550 1,410 1,410 1,320

P P P PP PP PP PP

1,570 1,470 1,370 1,960 1,760 1,670 1,570

1,420 1,320 1,220 1,760 1,570 1,470 1,370

1,930 1,680 1,670 1,600 1,620 1,560 1,520 1,470 1,420

1,710 1,480 1,480 1,420 1,420 1,370 1,320 1,280 1,230

DIN 4227 (GFR)

JIS G 3536-1971 (Japan)

opt — minimum tf0,2 — minimum tfip.c — minimum ö — minimum

5 . 2 - -6.0 7 . 0 - -9.5 1 0 . 0 - -13.0 1 . 5 - -3.0 3 . 0 - -4.9 3 . 0 - -7.5 4 . 0 - -10.0 2.9 3.5 4.0 4.5 5.0 6.0 7.0 8.0 9.0

tensile strength (MPa) yield strength (MPa) stress at 1 per cent tensile extension (MPa) ductility (per cent)

^0,2

0

^lp.c.

^(P.c.)

1,380 1,380 1,325 1,295 1,382 1,382

4 4 4 4 4 4

3.5 3.5 3.5 4.0 4.0 4.0 4.5 4,5 4.5

97 continued Standard

Diameter (mm)

Type

PN 22-178-73 (CSSR)

4 4.5 5 6 5

LR LR LR LR NR

WA BA NR LR P PP

#0,2

1,760 1,720 1,670 1,570 1,800

1,410 1,370 1,330 1,250 1,530

4 4 4 4 4

— wedge anchorage — button head anchorage — normal relaxation — low relaxation — patented — tempered

1000

Fig. 11-4. Relaxation of British tendons from the initial prestressing of 0.7 UTS at 20 °C 1 — stabilized wires; 2 — stabilized stranded cables; 3 — range of values for stress-relieved wires; 4 — alloy bars; 5 — range of values for stress relieved strands; 6 — range of values for 19-strand cables

— Patented high strength smooth steel wire, low temperature tempered, diameter 4.0 to 7.0 mm (Table 11-IV), coil weight 80 to 500 kg, internal diameter of coil 2,000 mm. — Patented, stabilized, high-strength smooth steel wire, diameter 4.0 to 7.0 mm, coil weight 80 to 500 kg, internal coil diameter 2,500 mm. — Patented, stabilized, high-strength steel wire with surface indentations, either untempered or low temperature tempered, diameter 4.0 to 7.0 mm, coil weight 80 to 500 kg, coil internal diameter 600 to 700 mm for untempered wire and 2,000 to 2,500 mm for tempered or stabilized wire. 11.3 STRANDS AND CABLES

In order to exploit to the full the mechanical properties of wires and maintain their flexibility while creating larger prestressing tendons, wires are stranded together into seven-wire, twelve-wire and nineteen-wire cables and occasionally into cables with more than nineteen wires. At present,

98 strands appear to be gaining popularity as tendon material. Cables used for this purpose should not contain any compressible components, such as a hemp core, nor must they be greased or oiled, as this causes problems with the effectiveness of grouting. Seven-wire strands have the simplest structure. Their load-bearing capacity is limited by the largest practicable diameter of the wire, viz. 6 to 7 mm; strands made from wires of greater diameter cannot be manufactured on the available equipment in the wire mills. Using 7 mm wire, a strand composed of 7 wires would have a diameter of 21 mm, and cross-sectional area of 269.25 mm 2 . Given a nominal wire strength of 1,400 MPa, the load-bearing capacity of such a strand would be 370 kN. Technical data for these strands are listed in Table 11-V. Untempered strands are supplied on minimum 600 mm diameter wooden reels and weigh about 2,000 kg. Low temperature tempered strands are supplied in coils weighing from 1,000 to 2,500 kg, and strands, composed of wires 4.0 to 6.0 mm in diameter with an overall diameter of 12 to 18 mm, are supplied in 3,500 kg coils. Another suitable cable structure is the 1 + 6 + 12 arrangement. In Czechoslovakia, for example, such cables are manufactured with a maximum diameter of 28 mm for anchoring in rock (these are unsuitable for prestressed concrete). The cable has a lower modulus of elasticity than cables of the 1 + 6 pattern, but has a regular layering and the internal layer does not slip during prestressing, once the cable has been locked at the head. General practice prefers the 7-wire strands also used in prestressed concrete. These strands are low temperature tempered and sometimes also stabilized, although, as stated before, the latter treatment is not essential for anchoring, considering the large permanent deformations which develop in rocks. In Great Britain and in some other countries, Dyform strands are used. These are strands which have already been through preliminary stressing and tempering (i.e. through the process of stabilization), and then are subjected to a compacting or dyforming process, whereby about 20 per cent, more of the nominal cross-sectional area is occupied by steel, compared with ordinary strand cables. Dyform strands have better mechanical properties and a 25 — 30 per cent, greater load-bearing capacity (see Table 11-V). In Great Britain, it is also possible to obtain 19-wire cables of the Seale or Warrington type (Fig. 11-5) with diameters from 22.2 to 31.8 mm. a) b) c) Fig. 11-5. Structures of cables used in Great Britain a) 7-strand cable, b) Seale cable, c) Warrington cable

3 PQ

o

J3

4>

E

'S

I g> cr

£ S

>3

· δ ^- S>

<

1 CO

o

©

q

©

fN fN fN fN

r- Γ^ r— r-

2 o vo Ü fN

©

o

O


Ό o CO

0

u

©

NO VO NO OO VO OO OO

q fN

r-' r«"

00

r-

ψ ^

"Ί

o o Ö

2 *-* *-*fN vo

co co O T * vo r- fN ON

fN

VO vo fN fN

VO vo vo vo vo vo VO en co co co co co co eo

vo vo

s

co fN fN fN N O O co fN

fN

^t *—i

TT ON 00 Γ^ Tt Tt oo' VO ON

fN co ON r- r© © fN r-»" co Ö vo ■
o _ „ Ö

o

o

i-H

fN vo

r- fN

vo OO NO ON CO ON fN ON VO ON co

ON OO fN

νθΛ

cr ,—Γ vo

«Ϊ

< ^ CO

NO

I-H

co co vo '--j ON -H'

s CO

<
3

O

V0 Tt

O

(0

U>

Ό

«-I

£

»i

B £ B

r- r- r- vo PC PC ti

o 8 o o ON o ON o «o ON

r^ r^ 1^

HJ

wo vo co fN

fN

s?s ~4

CO co

CO

Ö 00

-^ o o

CO

ON CS O O

vo vo NO VO

vo vo vo o vo vo NO vo 00

vo

co co

O

O o

fN

vo 00

ON

o

NO co fN

co vo ON co

fN

fN fN

o o

CO 00

vo 00 00

„

co O

ON

o

vo NO VO vo oo 00 OO OO

so OO

o

O

o

ON fN 00 vo O co OO fN OO NO O OO ON OO 1^ vo fN vo vo fN fN fN 00 vo CO co vo 00 ^ fN co fN CO CO

fN oo

vo

ON

fN vo

fN ON fN ^t fN fN ON CO vo vo CO ON CO

ON
vo r-

fN

_cö

c

«3

vo

2 fNoo O co O fN

CO

\ 0 (N

VO

60

S

CO CO 00 TJON fN

»o

O G PH

„ CO 00 Tf fN

Jo

ON O

^

00

fN

VO

r- fN q fN

-H

H

fN

q q q 9 P P «Ti CO VO vo r-' ON fN VO fN ^ H CN

q

fN r»# OO co ON

_; rt Γ^

Ö

CO ON VO fN fN VO ON

CO fN «O

0 0 VO 1—1 f N

PC PC tf c^ Pd P4 & ti >, >> >^ J J J

zzzz

q

co ON

fN

r-

00

vo vo «o «r> «Λ> VO vo vo co co CO co co CO co co

|

co co T-H

o

vo NO ON

t^

fN

»o fN

co

Tf OO

q q

q

ei

2 Ü

PQ

Ö r4 «n

ON

,_**ON r

O vo co vo
co CN VO

co ON

ON 1

NO co CO

PQ

fN

ON

ON

vo

t^ CO

CO

co vo VO fN fN

CO

CO vo vo ON

ON CO

vo co

ON

s

^

„

s

o o

vo rvo ^

OO NO VO

PC PC PC

z

1-H i-H

00 ^O »-H CO fN 00

vo vo vo co co*

CO*

fN

s

vo fN ON vo VO fN fN ^ H CN

ON

TJ-

5

vo CO VO

00 ON

vo «o vo vo fN V0

>υ

CO

z

r-

CO

vo fN

1^ fN

00 ON O N fN CO 00 co CO 00

o

O N 00 CO vo ON VO Γ*- vo vo 00

vo vo vo co CO

f-

PQ

00 00 OO

o o o o o o fN vo NO

fN

<

vo CO

vo v->

o vo

CO

3

vo vo 00 00 ON ON vo 00

o

vo

OO

00 ON ON fN 00 ON co

fN vo vo ON

vo

^

NO vo vo

o

ON

1

cm

CO 00 fN ON

Ü

VO CO CO vo CO

100

Large anchors are constructed using multiple cables which are arranged in the same way as the multi-wire cable. The manipulation of bundled cables is, however, more difficult than the handling of compound cables, and some contractors prefer compound strand cables for this reason. In Czechoslovakia, cables compound of 7 strands, each of 7 wires of 6 mm diameter, are used for 1.1 MN anchors, and for 4.0 MN anchors compound rope (strand) cables Hercules are used, these comprising 37 strands each of 19 wires of 2.9 mm diameter, giving a total cross-sectional area of 4,320 mm 2 . The cable diameter is 100 mm, the weight is 34 kg/r.m., and the nominal load bearing capacity is 7.6 MN (Fig. 11-6).

Fig. 11-6. Hercules cables; 39 bundles of 19 wires of 2.8 mm dia. nominal load-bearing capacity 7.6 MN

The disadvantage of stranded cables compared with multi-wire cables is their larger tendency to creep and their lower tensile strength. This creep, however, is still small in comparison with that of the rock, and does not detract frcm the advantages of stranded cables for anchoring. The difference between the actual cable strength and the nominal cable strength (the sum of the strengths of the individual wires) arises from the stress distribution in the loaded cable. Its magnitude depends on the structure of the cable, the extent of transference of wire loading, and the extent of defects in cable manufacture, etc. Any drop in the actual cable strength below its nominal strength depends on the type of structure involved. It amounts to 13 %, for example, in compound strand cables. For an anchorage consisting of patented 3.6 or 4.1 mm diameter wires, the Grands Travaux de Marseille contracting company specify a strength range of 1,700 to 2,100 MPa and a load-carrying capacity of at least 90 per cent, of the sum of the individual wire bearing capacities. The relaxation after 100 hours under load should be less than 5 %. A lower actual strength need not be an obstacle to use of stranded cables for anchoring; this disadvantage must be balanced against the greater ease of handling and transport on site, the greater cost effectiveness, easier fixing,

101

and improved resistance to corrosion. This last advantage has been demonstrated at the Vir Dam in Czechoslovakia, where after 10 years of service, cables composed of straight wires were found to be corroded along the sections passing through the open air, while stranded cables remained intact under the same adverse conditions. The disadvantage of steel rope cables and compound strand cables is that specially designed strong anchoring heads must be used where the cable is attached to the anchored structure (see Chapter 16).

Chapter 12 P R E P A R A T I O N OF A N C H O R S

The preparation of modern anchors, particularly with respect to anticorrosive protection, requires specialized equipment. Special workshops are assigned to this work, which is entrusted to highly skilled men. The equipment and techniques used in these workshops vary according to the tendon material that is being handled.

12.1 PREPARATION OF BAR ANCHORS

Bar anchors are the simplest to prepare. The required length of bar is cut and a thread is cold rolled at the head end to take the fixing nut. The thread should not be cut, because this reduces the effective cross-section of the bar. At the root end of the anchor, threading improves the bond between the anchor and the cement (grout or resin) used to fill the borehole in this section, or serves for the fixing of the anchor base (Fig. 12-1). The bars are supplied in 12 m lengths for ease of transport and handling. Longer anchor bars are made up by connecting these bars with couplings (Fig. 12-2).

Fig. 12-2. Coupling of prestressing bars

103

The preparation of anchors from bars with rolled-on threads is extremely simple. The preparation can be carried on directly on site, including the application of an insulating coating or the slipping-on of a plain insulating pipe over the free section of the bar corresponding to the anchor tendon. Suspended pipes, which provide double anticorrosive protection along the anchor root (see Chapter 18) are preferably mounted and filled with grout or resin in the shop where special injecting equipment can be used and careful control exercised. Bar anchors are sometimes inserted into, and fixed in pipes crimped along their entire length, in order to increase anticorrosive protection. In these cases the root section is separated by a seal which must also be made in the workshop. 12.2 CONSTRUCTION OF MULTI-WIRE ANCHORS

The assembly of large cables from individual wires is time-consuming, and requires a large work area. The wires are straightened and cut to the required lengths, then they are arranged in layers with the aid of a metal sheet guide (Fig. 12-3) with holes for the wires, and bound together every 0.5 to 1.0 m. The most important part of this work is the positioning of the wires, because their cross-sectional area is not fully utilized and their strength is reduced if the wires are not mutually parallel throughout. Thus, for example, the 1 MN anchors, that are used in Czechoslovakia (24 wires of P 7 m m diameter) are constructed in two layers (Fig. 12-4). The internal layer consists of 8 wires, each 7 mm in diameter, around a helix of internal diameter 12 mm and axial length per coil turn 5 to 6 mm. Around each layer is wound a 2.2 mm diameter wire. The winding of this wire around the internal layer serves as a spacer between the two layers, and together with the central helix, ensures reliable interpenetration of the cable by the grout.

Fig. 12-3. Putting together cables of patented wires P 5 mm dia

104

Fig. 12-4. Anchor head and cable made from 24 P 7 mm diameter wires (Hlasivec—Michälek system) 1 — central helix, 2 — internal layer (8 wires), 3 — external layer ("16 wires), a) — head, b) — cone, c) — wires

For the assembly of a large number of anchors, a mobile workshop may be set up. The cable is assembled on a special piece of equipment consisting of a carriage fitted with a perforated guide, an inserter for the central helix, an end gauge, winding equipment with storage space for winding wire, and equipment for the application of an anticorrosive coating. The straightening, positioning and winding of the wire are fully mechanized. A new system for the manufacture of tendons has been developed in the USSR. A continuous wire, or a bundle of several wires, is stretched around two pins spaced at a distance equal to the length of the cable required. A special carriage is used for the winding. In other countries, anchors composed of straight wire assembled according to the BBRVsystem have been used over the last few years (Fig. 12-5). They can accommodate anchoring forces of up to 12 MN, providing a sufficient number of wires is used. The setting of the wires in large bundles is facilitated by using a base, which consists of a steel plate with holes through which the wires are inserted; the ends of the wires are then hammered to form small heads (see Fig. 16-5). The diameter of the wire may range from 4 to 12 mm according to the manufacturer's data. A cable composed of 121 wires of 7 mm diameter, with a strength of 7.7 MN, may be cited as an example. The wires of large BBRV cables are not positioned by means of spacing grids as this would greatly increase the cable diameter; thus sometimes the

105

Fig. 12-5. BBR V anchors used in anchoring the Spullersee Dams 1 — grouting pipe, 2 — anchor head, 3 — supporting ring, 4 — load distribution plate, 5 — reinforcing coil, 6 — sheath into which the head is inserted prior to the stressing of the cable, 7 — sealing collar, 8 — anchor base, 9 — de-aeration pipe

only spacing is the helical coil at the core of the cable. The wires are automatically measured and cut to the required length on a special workshop table. When the wires are arranged together, the bundle is straightened by pulling the wires through a guide and a base piece where the ends are cold hammered to form heads (Fig. 12-6). To keep the wires parallel, the cable is held by the base while the guide is pulled in the opposite direction. Finally, the cable is twisted three or four turns to take up any slack in the wires arising from lack of straightness of the axis and it is then bound around with a thin wire

Fig. 12.6. Assembling BBRV cables in the factory

106

in sections 3 m long. The twisting of the cable also allows it to be bent or wound on to drums for transport to the site. Several other systems of anchoring cables also employ straight wires; The Polensky & Zöllner system used in the GFR and Austria (see Fig. 16-3) and the Losinger system (VSL) of Switzerland (Fig. 12-7) are examples. It should be noted that some of the cable assemblies mentioned here are at present going out of use and are being replaced by other types developed by the same companies, but using steel ropes as the basic anchor ebment (see Fig. 13-26).

Fig. 12.7. Cable anchors 2 MN prepared for transport (Ingstav Brno) 1 — flexible sheathing tube, 2 — grouting pipe, 3 — locking ring, 4 — spacing ring; A — spacing of rings 60 cm, B — length of fixing section 3 to 5 m

12.3 C O N S T R U C T I O N OF ANCHORS FROM STRANDS

Strand anchors are prepared in the same workshops as those in which anchors composed of individual wires are made. The process is less timeconsuming because a smaller number of tendon components has to be manipulated. The use of ropes is a sign of the growing tendency to industrialize the manufacturing process, that is, to transfer cable construction from the anchoring site or temporary workshop to specially equipped factories. (Fig. 12-8). The strands are supplied to the works on reels or as reelless packs The coils are arranged on pins in such a way that the prescribed number of ropes required for the cable can be unwound from them along the shortest possible distances. An important part of the fabrication of cables is the provision of insulation which must give durable anticorrosive protection. The insulating layers are often formed on the individual ropes before

107

m-

is

**; :JftS** -V* * ^

*%,? Ht, '^'

ΜΧΆ^

Fig. 12-8. Anchors system Losinger iVSL)

°)

Fig. 12-9. Tendon protection a) — unprotected strand, b) — plastics covered strand, c) — greased and plastics covered strand, 1 — plastics covering, 2 — protective grease

the latter are incorporated into the cable, by coating and pressing the insulating compound on to the surface of the rope (see Chapter 18), or pulling the ropes into plain or crimped pipes. The insulation type depends on the function of that part of the anchor being treated and its overall structural arrangement. This work, in particular the coating of insulation, is also now being transferred to special factories. The bundles of strands constituting anchors of high load-bearing capacity are usually inserted into plastic pipes (Fig. 12-9). The workshop, in this case, must be equipped for the filling of the internal free volume of the pipe with an insulating compound; the latter may be one of various vaselines

108 Fig. 12-10.. Transport of tendons in loops

(see Chapter 18) or a mixture of asphalt and synthetic resin. In the grouted section, that is, in the section sheathed with a crimped pipe, the pipe is internally filled with grout. In some cases the cables, instead of being inserted into pipes, are wrapped with a textile bandage impregnated with insulating vaseline. The bandage is wound on so as to be self-overlapping, and if need be, covered with another layer of plastic material to protect the anchor from mechanical injury during manipulation. The bandage is put on with special winding equipment. The cables are prepared in specified lengths and delivered to the site in bundles (Fig. 12-10), or in reels or as reelless packs (see Fig. 12-8).

Chapter 13 F I X I N G OF A N C H O R S IN ROCK AND S O I L

The fixing of anchors so that the tension within them is taken up by the ground is achieved by three basic methods: — by mechanically bracing the anchor foot (base) up against the rock at the end of the borehole; — by bonding the anchor tendon to the rock or soil with cement; — by fitting an expanded base (i.e. a bulb) at the distal end of the anchor. The face of the bulb abuts against ground when the anchor is prestressed. The first of these methods relies on the forces of friction set up between the steel jaws, or other parts of the base, and the borehole walls (see Fig. 9-7). The ground is subjected to considerable radial pressure concentrated over the small area of contact with the base. The second method relies upon the bond strength developed between the anchor root and the rock or soil of the borehole wall (see Fig. 9-8). This bond is achieved by using a suitable cement (cement grout, synthetic resin). The rock is mostly stressed by shear forces which are distributed over the relatively large lateral surface area of the root cylinder. The third method depends upon the strength of the ground, or its resistance to the extraction of the deep-situated base (or bulb). The rock in this case is subjected to a locally concentrated compression. In some special cases, combinations of these fixing methods may be used. The fixing method that is finally selected depends on the mechanical characteristics of the rock or soil, the magnitude of the tensile force for which the anchor has been designed, the design of the anchor itself, and frequently also the facilities and equipment available to the contractor.

13.1 MECHANICAL FIXING OF ANCHORS

Mechanical fixing is mainly employed in temporary short anchors (bolts) in strong rocks. The mechanical fixing device (the fixing base) is fastened at the end of the anchor inserted into the borehole, and the fixing is achieved by the expansion of this base against the borehole wall; in theory the pressure of expansion may be increased until the compressive strength limit of the rock is reached. When the anchor is under stress, displacement of the anchor base is prevented by the friction of contact with the borehole wall. The

110

<0

Fig. 13-1. Fixing of a wedge bolt in a borehole [113] a) — wings of the cleft forced against the rock of the borehole wall at the fixing position, b)y c) — parabolic shape of area of contact between the bolt wing and rock on which stress is concentrated, (A) after loading the bolt with tension, d) — magnified view of the wing surface at the contact area. Steel in the middle {A—B) is damaged by the shear stress, and crushed rock is pressed into the cavities. Edges become knurled

Ill

resistance to extraction of the bolt is determined not only by the magnitude of the friction, but also to a large extent by the shear strength of the rock against which the base is forced (Fig. 13-1). The stress state in the rock caused by the pressure of the mechanical base has been investigated by using photoelastic measurement and computing techniques. For example, Ewoldsen [65] used the finite element method. to investigate the axial, radial, and tangential stress components around a bolt fixed into elastic homogenous rock (Fig. 13-2). He found by computation that if the anchor base were given a prestressing force of 44 kN (with borehole fixing section 51 mm diameter, 76 mm long), a maximum compressive stress of over 7.6 MPa was created in the rock, parallel to the anchor, at the point where the base was connected to the tendon; at the other end of the base there was a nearly identical tensile stress in the rock. Because of this stress, cracks can appear behind the base, near the borehole bottom. When this occurs, however, the pattern of rock stress does not change significantly; indeed the distribution of stress may become more uniform, as can be seen in the bottom diagram of Fig. 13-3. The distribution of the radial and tangential stresses in the close vicinity of the mechanical fixing is also shown. Similarly, the stresses in the rock have been assessed for a 3 m-long bolt a) axial

b) tangential

c) radial rock surface

L

0

i

1

I

J

.

m

i

I

0

.

i

1

i

i

2m

I

o

.

1

1

.

1 —

2m

Fig. 13-2. Stress state of rock in the vicinity of a bolt with a prestressing force of 44 kN

5 tress

( stresses in psi\ pressure marked with positive sign ) axial 6a

tangential 6t

raaial

ar

Λ?

scale of lengths:

-50

0

5

ID

15 cm

conversion table ■■ D si

50

too

200 300

too

500 600 700 800 900 1000 1100

,'
%

21 28 35 U2 L9 55 63 70 77

OM 0.66 1.37 2.06 2.75 3M U2 4.80 550 6.16 6.85 i Z55

hissure

Fig. 13-3. Stresses in the vicinity of an anchor base mechanically expanded against the walls of the borehole by a force of 56 kN, and pulled by a tension of 44 kN (according to H. Ewoldsen)

113

under a prestressing of 44 kN. The iso-stress lines in Fig. 13-2 show a markedly steep decline in stress with distance from the borehole. At a distance of 10 cm, the radial and tangential stresses amount to less than a twentieth of the maximum value, and they practically disappsar at a distance of one metre. The compressive stress along the borehole (between the supporting washer and the base) also declines rapidly towards the centre of the zone to a figure as low as one thousandth of the maximum value calculated at the base front. By analysing the forces acting within the base and in the contact area between the base and the rock, formulae can be derived for calculating both the optimum fixing force for a mechanical base, and the fixing strength of a given rock type. These problems have been discussed in detail by A. Hugon and A. Costes [94], and by T. A. Lang [113]. Tensile tests carried out on mechanical fixings in rock do not agree completely with results obtained using these formulae. Differences of up to 100 per cent are attributable, above all, to highly inaccurate values being substituted in the formulae. Thus, for example, the rock strength at the fixing point of the borehole wall is determined from laboratory strength on rock samples, and the value for the coefficient of friction between the steel base and the rock is simply an estimate. Testing of the load-bearing capacity of bolts on site is comparatively easy, and it is always better to rely on values for the average fixing strength of a given type of anchor and rock, obtained from in situ tensile tests. Mechanically fixed anchors have the following characteristic properties: — They acquire full strength immediately on fixing, which means they can be prestressed or loaded as soon as they have been placed. — The fixing against the walls of the borehole is usually in a short section of 10 to 20 cm. This introduces large stresses into the rock, so that the demands on the rock strength at the fixing point are very considerable. — The fixing strength is principally determined by the fixing force with which the jaws of the mechanical base are pressed against the rock. — When a mechanical anchor is overloaded, the expanded base slips along the borehole walls if these are of hard rock, whereas in softer ground the base may be pressed deep into the borehole walls, and therefore breaks under overload. — They are not usually protected against corrosion and therefore their life expectancy is limited. Because of this they are mainly used for the temporary securing of rocks. — Their cost is greater than that of bonded bolts, because their manufacture is more complicated and they are usually made of high quality steel; the borehole diameter must also be greater (30 to 70 mm). Mechanical bases differ according to the method of expansion in the borehole.

114

13.1.1. Fixing by thrust (wedge base) The anchor with a wedge base is the oldest type, and the simplest to manufacture. The tendon of the anchor, which is made of circular crosssection steel of diameter 22 to 26 mm, is given a longitudinal cut at the basal end, and a steel wedge is inserted into the cleft thus formed. The bolt is fixed by hammering inwards so that the wedge is forced into the cleft which thus opens and presses against the borehole wall (Fig. 13-4). The wedge-base bolt has been used for many years, particularly in mines. This type of bolt can also be made quite easily on site. The cleft is cut with a welding torch, and the wedge is forged from a piece of drill rod or similar material. If a wedge bolt is to be fixed successfully, the optimum dimensions and correct positioning in suitably hard rock must be observed. An optimum specification for such a bolt is as follows: diameter of bar, 22 to 26 mm; overall length, up to 3 m; length of cleft and wedge, approximately 14 cm; taper of wedge 7°; difference in diameter between bar and borehole, 5 to 7 mm maximum; the length of the borehole should be 10 to 15 cm less than the bar length. When fixing the anchor, deep wedging of the cleft at the borehole bottom must be achieved, and the best tool for this operation is a medium sized pneumatic pick (possibly with a pneumatic strut) fitted with a cut-away bit and guide pipe (see Fig. 13-4). Driving the bolt in with a mallet generally has little effect. The wedge-base bolt can be used successfully in all hard rocks. Loadbearing capacities of 70 to 120 kN can be attained, with a partial extraction of up to 10 mm. A correctly fixed wedge-base bolt did not show any decrease in fixing strength during repeated tensile tests, even after several years. 13.1.2 Fixing by tension (tensile base) Bolts with tensile bases are the simplest both in terms of fixing and prestressing. Following insertion into the borehole, the bolt can be loaded immediately by pulling in the cuneiform or conical support located at the end of the bar between the expanding toothed jaws. This is usually done at first by hand and then by a nut at the external end of the bolt bar. Bolts of this type are made in various countries. The best known is the Goldenberg bolt made in France (Fig. 13-5). Another arrangement of this type is the GD anchor. The conical end of the bar is pulled into four expanding jaws made of high quality synthetic material (Fig. 13-6). This material has a greater plasticity and lower modulus than that of steel, but shows a similar modulus to that of rock. Thus the transfer of force from the steel tendon to the rock face is more effective —sometimes more so than that achieved with a steel base.

10

Fig. 13-4. Wedge bolt and method of driving-in using a pneumatic pick

l|

iw

ψ

^

1800

20 cm

1

HO

*

20

7

~^ί$0\

ί

t

116

A special type of bolt fixed by tension has been developed by the American Worley Co. (Fig. 13-7). By tightening the nut on the external end of the rod, both parts of the bolt are forced against the rock along their entire length. The bolt may thus be used in weaker rocks. If desired it may be taken apart and used again.

Fig. 13-5. Tensile anchor base of Goldenberg bolt

Fig. 13-6. Synthetic base of GD anchor (Farex-Sv/eden); 4 jaws surround the conical bar end

13.1.3 Fixing by screwing (threaded base) Bolts with threaded bases are fixed by turning the bolt bar. In this way a conical or cuneiform piece at the distal end of the bar is drawn between the jaws which are bridged by a holder (Fig. 13-8), and the jaws are thus forced against the borehole walls. Penetration of the rock when the jaws can be opened no further is in practice between 3 and 10 mm. The softer the rocks the deeper is the penetration. The rolled 4 mm thread on the bar end allows rapid tightening with high torque transmission. The initial locking of the bolt while it is under tension may be carried out by hand turning using a 50 cm lever. A torsional moment of 350 Nm induces a prestress of about 70 kN in the bolt. This type of fixing is somewhat more complicated in its manufacture and therefore is also more expensive than the types mentioned above. It is, however, generally considered to be the most reliable, because the base can be expanded to a greater width, and requires only light initial contact between the expanding components and the borehole walls. It also evens out more effectively the accidental roughness of the borehole circumference, thus making contact with areas of higher compressibility in the borehole walls.

117

Fig. 13-7. Worley bolt (USA) with anchoring effect along entire borehole length

Fig. 13-8. Threaded anchor base of Pat tin bolt

The contact surface area and dimensions of the expanding components may be adjusted according to the rock strength. For hard rock, a small contact area and shallow, but sharp, indentations on the surface suffices (see Figs. 13-8, 13-12). For medium hard and soft rock, the expansion surface requires a larger contact area and a rough face (see Fig. 13-10); in some cases two or more co-axial shells are used. The load-bearing capacity

118

of such bolts can range from 170 to 220 kN in strong rock. Their ultimate bearing capacity, however, is usually determined by the smallest diameter of the bolt rod (16 to 36 mm). When the bolt is no longer required in the same position, the rod may be screwed out of the shell, extracted from the borehole and used again. Sometimes, by giving repeated knocks to the rod just before it is completely unscrewed, it is possible to retrieve also the expanding base from the borehole. The possibility of recovering bolts is only of value in mining. Mechanically fixed bolts with threaded bases are much used nowadays, and they are made by many manufacturers all over the world. Renowned makes, such as Pattin (see Fig. 13-8), Ancrall (Fig. 13-9), Lenoir et Mernier in France (Fig. 13-10), Rawlbolts, Bayliss (Fig. 13-11), Torque Tension in England, Dywidag in the GFR, Bail in South Africa (Fig. 13-12), Williams

b) Fig. 13-9. Threaded anchor base of Ancrall bolt a) — dismantled, b) — assembled

Fig. 13-10. Expansion bolts made by Lenoir et Mernier

119 Fig. 13-11. Threaded anchor base of Bayliss bolt (England). Above: as inserted into the borehole; below: after the sleeve is opened by pulling-in the cone

^IÄMA! Fig. 13-12. Threaded anchor base of Bail bolt (Republic of South Africa)

in the USA, Titan in Australia, may be mentioned as examples. A large variety of bolts with expanding bases is offered by the French firm of Lenoir et Mernier Co. They manufacture 8 types of base with different dimensions (Table 13-1), for various types of rock. Most of the bases are fitted with a special pre-bolting spring, which provides for the initial expansion on insertion of the bolt into the borehole. TABLE 13-1 Mechanical bolt types made by Lenoir et Mernier Type

Initial diameter [mm]

Expansion [mm]

Length [mm]

Surface area of contact [cm2]

Borehole diameter tolerances [mm]

Rock

31 34 36 41 41 46

LN LS UM UP UM UP

31 34 36 41 41 46

+ + + + + +

16 16 18 18 18 18

98 75 120 100 120 120

80 70 100 100 140 180

close fit close fit 36+4 41+5 41+5 46 + 8

56 UM

56

+ 18

200

300

56 + 8

66 UM

66

+22

250

400

66 + 9

hard fairly hard fairly soft soft to hard compressible all types of rock all types of rock all types of rock

120

13.1.4 Fixing of hollow bolts by expansion (friction bolts) These bolts, made from steel tubes, are pressed against the borehole wall along their entire length by expansion of the tubes. The resistance against any displacement of this anchoring tube fixed in the rock is due to friction between the rock and tube throughout its full length. Bolts of this type are usually indicated as friction bolts [188c]. A special type of bolt, fixed in the borehole along its entire length of 2 m by swelling of a steel tube by means of a high water-pressure was presented as Swellex by the Swedish Atlas Copco (Fig. 13-13). Swellex is manufactured from a steel tube of 41 mm diameter, which has been mechanically reshaped to assume an outer diameter of only 28 mm. Sleeves are pressed onto the ends sealed through welding. One of the sleeves retains a washer in place and has also a small hole to allow water to be b)

o)

Fig. 13-13. Swellex bolt a) — bolting tube before expansion, b) — bolting tube fixed in the borehole by expansion

121

injected into the tube during a quick installation of the bolt in the borehole. High-pressure water expands the bolt immediately. As the borehole diameter (30 to 39 mm) is smaller than the original diameter of the tube, a tongue is left inside the profile (see Fig. 13-136). This tongue acts as a spring when the water pressure is released, and produces radial forces which continue to press the bolt against the rock walls of the hole. During the swelling process, the length of the bolt is reduced due to contraction, resulting in forcing the washer of the bolt against the rock face. About 50 bolts may be fixed in this way in an hour's time. A steel tube provided with a slot throughout its length, which is forced into a borehole of a corespondingly smaller diameter, has a similar, although somewhat smaller, effect. This type of friction bolt is currently produced under the trade name Split Set Stabilizer by the American Ingersoil—Rand Co according to the design of J. J. Scott [188] (Fig. 13-14). This bolt is driven home into a borehole of 35 mm (13/8 in.) diameter by means of a percussive or vibrating equipment as jackdrill or stopper, but air or hydraulic drifter or rotary roof bolter can be used as well. When set, the initial holding power is about 27 kN and further increases with time. When the pull-out force acting in the roof on the bolt plate exceeds the bolts holding force, the tube slips a little while maintaining its friction resistance. This possibility of support yielding without failure is preferable for rock mass stabilization as mentioned in the next Section.

Fig. 13-14. Split set stabilizers of Ingersoil—Rand Co. a) — view of the bolting tubes with rings and washers,/?) — cross-section of the slotted tube

122

Another advantage of these hollow bolts is that they allow water drainage of the supported rock. Split Set bolts are delivered in lengths varying from 0.9 m to 2.44 m (3 ft. to 8 ft.) in 0.3 m (1 ft.) increments. Because of their simple installation technique they are widely used, especially in the United States metal mining industry, where they form approximately 50% of all roof fixtures [188a]. 13.1.5 Controlled yielding bolts Ordinary rock bolts yield and extend up to 18% and then they break [188c]. Recent knowledge in rock mechanics concerning the advantages of allowing some yielding of the supports in underground caverns, has led to the design of bolts with controlled yielding [61, 152]. A special coupling allows controlled elongation of the bolt if its load exceeds the admissible limit. This elongation of the bolts which are used to reinforce caverns, takes place automatically and allows the rock to undergo slight movement. This movement reduces the pressures in the rock and the cavern develops a new state of equilibrium. The elongation coupling of the bolt developed in the USA (Fig. 13-15) [61] consists of a sleeve around the outer part of the stem of a standard bolt, the latter being provided with rolled threads with an overall diameter greater than that of the stem. Because the inside diameter of the inner end of the sleeve is less than th? outside diameter of the threads on the bolt, the bolt

<0ο°Λ0Φή°°* 'ο^ο 0- ο0Ό ρ;νΟ

ν

0

^ mine roof ;Ό:.'00·Ο ■ ο or other •■\J'd'i'-oQ \γ. wall rock.: ■:· %0o

stem portion of standard . v rock bolt

as

\c o ° O O

.

rOD

0° o note rolled.■*■<■< 7 ^ 7 threads .ο°(/* ο1 OU.-O-L'-G

Η-0ΐ

Op o<-. t\

Q\ •■'o°Z.oc[
0 Vr C °■ ° t C X Cn( t-2— sleeve , C P ; y^

plate

-nut

Fig. 13-15. Controlled yielding coupling for standard rock bolts developed in the USA

123

may be prestressed to the permitted load by tightening the nut threaded on the external section of the sleeve. When the loading of the bolt reaches a predetermined limit value, the threads on the bolt stem are partially stripped with the stem moving into the sleeve to some extent. In this way, tension in the bolt automatically drops below the admissible limit, which is determined by the strength of the threads. This arrangement allows the bolt to elongate while it continues to exert a strong restraining force on the surrounding rock. Yielding bolts of a different design are offered by the Austrian Meynadier Company. A special Meypo head is slipped onto an anchoring steel rebar with a diameter of 24 mm without thread at its end; this head induces an effective thrust of a headplate on the rock and at the same time allows a yielding of up to 50 cm while transferring a constant tensile force of about 214 kN. Smooth yielding is provided by means of steel balls filling out the inner conical cavity of the head, closed by a nut.

13.2 FIXING OF ANCHORS WITH CEMENT

The bonding of cement (cement grout, cement slurry, synthetic resins) to both the anchor tendon and the ground is the most often utilized method of fixing anchors into rock and soil. The cementing of an anchor into a borehole is carried out over a comparatively long section of the anchor (1 to 10 m); the specific load on the rock or soil is therefore small. This method of fixing requiring a long root is particularly effective in weak rocks and loose soils; it is also used for transferring large tensile forces into strong rocks, and may be used with success even in cohesive clayey soils. The anchorage is protected from corrosion in the root section, but it can be loaded only when the cement has hardened. The efficiency of the fixing depends on the adhesion of the cement both to the anchor surface and to the rock or soil in the borehole. 13.2.1 Design of anchors fixed with cement The length and cross section of the part of the root cemented in the ground is usually determined under the assumption that there is uniform cohesion of the cement over the area of contact, both with the steel tendon and with the rock or soil of the borehole wall. The contact area, and consequently also the root dimensions, are determined from the simple relationship between the tensile force, P, and the average shear strength at the surface of the cemented root, where the latter is considered to be of regular cylindrical shape:

124

m . P — π . d2 . / . tb, or between the force, P, and the bond strength between the cement and the surface of the steel components of the anchor root section: m . P = n . π . dl . / . τα, where m dl d2 n / xb xa

= safety factor (1.5 — 2) — diameter of the cemented section of the anchor tendon = borehole diameter = number of steel components in the root = fixing length of the anchor root = bond strength between cement and rock or soil (working value) = bond strength between cement and steel.

In designing the fixing the weakest component is considered, and means are found for increasing the load-bearing capacity of the anchor solely by improving the function of this component. A sufficiently large fixing area is established first of all, as determined by the length and diameter of the anchor root. The most frequently used cement for fixing either short or long anchors in the ground is grout. Information concerning its adhesion to various types of rock, soil, and the steel of the anchor root, is contained in the following Sections. IS.2.1.1 Cohesion between grout and rock The bond between the grout and the borehole wall depends on the strength of both the rock and the grout, the roughness of the borehole walls, the cleanness of the borehole, and the area of contact between the grout and the rock— a factor partly depending on the length of the root. With increasing root length the required strength of the bond decreases in proportion (see Chapter 9.3.2 and Fig. 9-9). The approximate strength of the grout-to-rock bond is ascertained in the laboratory, using cores from anchoring boreholes or from exploratory boreholes. The core is placed inside a strong steel mould (Fig. 13-16) which is filled with grout to a predetermined level. After at least seven days for hardening, the core is pressed out of the mould in a laboratory press. Bond strengths obtained in this way are usually somewhat higher than those found in field tests because the contact area in the laboratory test is much smaller, but particularly because the core expands under compression rather than contracting, as it would under tension. A test in which the core were under tension would correspond more accurately to the mode of stressing of rock

125

surrounding a tensioned anchor. The adopted working strength of the bond is therefore set by introducing a high safety factor, 3 or 4. According to Littlejohn and Bruce [120], the ultimate cohesion between cement tnid rock may be taken as being approximately one tenth of the compressive strength of the rock, up to the strength of the hardened grout. ^ΤΗΓ^ ^^^^|

Fig. 13-16. Laboratory cohesion test between rock (drilling core, F) and grout (documentation of Losinger Co.)

Assuming that the strength of grout is 42 MPa, the ultimate bond strength is then 4.2 MPa, and the admissible bond strength is 1.4 MPa, taking a safety factor of 3. Coates [32] refers to a maximum working value of 2.45 MPa for the bond strength, but with a safety factor of 1.75; hence the ultimate strength is the same (4.3 MPa). For weaker and partly weathered rocks it is more correct to take the ultimate bond strength as being 20 to 35 per cent of the compressive strength of the rock [120]. Koch [120] suggests working values for cohesion of between 0.35 and 0.70 MPa for soft rocks, 0.70 and 1.05 MPa for medium hard rocks, and 1.05 and 1.4 MPa for very hard rocks. According to the Australian Standard, a value of 1.05 MPa is satisfactory for the majority of strong rocks. Weathering of rocks significantly reduces their bonding strength with grout. The most reliable measurements of bond strength are obtained in pull out tests at the anchoring site. The average value for the ultimate bearing capacity of the anchor divided by a safety factor of 1.5 to 3.5, is used in the design of the anchorage. A lower safety factor may be used for compact, strong rocks and for temporary anchors, while a higher safety factor should be applied for weaker, jointed, or weathered rocks, and for permanent anchors. Table 13-11 lists bond values obtained in field tests for different types of rock at different sites [120]. The ultimate values could not be obtained in all cases. It appeared that even for similar rock types the bond values varied considerablv. This variation arose from local deviations and ir-

126

regularities in the geological structure of the rock mass, from the different design details of the test anchors, from different drilling and flushing methods as well as from different pressures used in the grouting of the test anchor roots. TABLE 13-11 Rock/grout bond values found in field tests [120] Rock type

Test bond strength [MPa]

Granite Basalt Sandstone

0.81, 1.24, 1.72, 1.72 0.72, 1.0, 3.6 0.4, 0.84, 0.95, 0.98 1.2, 1.56, 1.58 1.42 0.63 1.19 0.55 1.66 0.88 0.21,0.28,0.36

Limestone Limestone (loamy) Limestone (fissured) Limestone (with marly bands) Dolomite Mudstone Marl Shale Shale and sandstone Shale (strongly weathered) Chalk Quartzite Breccia Slate Slate and greywacke Micaschist Micaschist (very sound) Micaschist and D/;ite-gneiss Meta tuff (weathered)

Ultimate bond strength [MPa]

6.37 1.73 2.83, 4.56, 4.80

1.8

0.63 0.10 0.39 0.7 1.02, 1.32, 1.72 0.93 0.90, 1.24 1.4 0.65, 0.80, 0.92 2.16 0.80 0.29

1.80

The stress distribution along a cylindrical anchor root was studied theoretically by the finite element method. [32] The stress distribution depends on the ratio of the elastic moduli of the anchor material and the rock, respectively. The smaller the difference between these moduli, the greater is the stress concentration at the proximal fixing end of the anchor. The greater the difference, the more the stress diagram approaches uniformity of distribution. For a modulus ratio higher than 10 (i.e. for soft rocks with an elasticity modulus of less than 2,000 MPa), the stress distribution is uniform with respect to most of the anchor root length. For stronger rocks,

127

on the other hand, the stress concentrated at the proximal end of the anchor is five to ten times greater than the mean theoretical stress. Thus, under high loading of the anchor partial debonding takes place at the proximal end, and this debonding progresses towards the distal anchor end as the load is further increased.

W

, . „„0 strait L / . .

?

Is

*.0 131mm ,

ψηψ r-|

V (— —1 -

3.0 1

1

S3

«ί;

2.0

um/

//

1.5m/

,. y

1

1 1

1.8SMN

anchor cross-section 5V No. wires 7 mm diameter

Fig. 13-17. Strain distribution along the tendon in the fixing zone of a 2.2 MN capacity test anchor [144]

Non-uniform stress distribution along a long grouted anchor root transferring tensile forces to the rock has been confirmed in experimental tests carried out in situ by various authors [16, 144, 99a]. By means of strain gauges fixed to the anchor tendon at various points along the fixing section, the decrease in tension from the proximal to the distal end of the root may be observed, together with the changing degree of transfer, of tensile force from the anchor to the rock. Results obtained for a BBRVanchor with a bearing capacity of 2.2 MN, length 8 m, are shown in Fig. 13-17 [144], Thus under a load of 0.5 MN, the force was transferred at the upper end of the root with an average bond/rock stress of 0.22 MPa. When the load was increased to 1.85 MN, the bond failed along an upper root section of length 3.9 m, and the load was taken up by the remaining lower root section (4.10 m long) with an average bond stress of 0.98 MPa. The anchor base withstood about 0.3 MN. At a loading of 2.8 MN, a comparison between the theoretical and measured extensions of the anchor tendon in the fixing section suggested that the tendon was debonded along its entire length so that the load was transferred completely to the anchor base. Nevertheless, the ultimate loadbearing capacity of the anchor had not been reached. The average value of the bond strength computed with the assumption of a uniform load distribution along the root, was 0.65 MPa; this was much less than the actual value obtained with a test load of 1.85 MN, and was also much lower than the shear strength of the grout. The results of the above-described experiment are very instructive. They prove that in anchorages in strong rocks, the grout/steel bond is the weak point of the fixing and they also confirm that the most efficient fixing is achieved with an anchor base, even if the base is not formed in an expanded anchor borehole.

128

13.2.1.2 Cohesion between grout and soil Where anchoring is carried out in soil, the weak point of the fixing is the interface between the anchor root surface and the soil. The resistance to extraction of the anchor root from the soil is determined by the shear strength (force of friction) at the soil/root interface; this value may be expressed by Coulomb's relation: τ = σ . tg φ + c. In non-cohesive soils the grout/soil bond strength at this interface is usually larger than the shear strength of the soil, and it is possible to determine the shear strength around the root using the parameters obtained for soil, and derive the soil stress, σ, on the root, from the weight, of the anchor root overburden. Tensile tests of anchors fixed in soil have indicated that the resistance to extraction is usually higher than that calculated from the ultimate shear stress around the root. L. Hobst explains this as being a result of the fact that the overall diameter of the root, d2, delimited by the soil/root interface at which the shear strength is critical, is greater than the diameter of the borehole itself. Moreover, over the entire surface of the root which has an irregular form owing to uneven penetration of the grout, a transverse pressure appears which considerably increases the effect of the frictional component of the fixing strength of the anchor in the soil. Hobst used this phenomenon as the basis of a formula whereby the minimum necessary embedding depth of the anchor may be calculated (see Section 10.3.1); the validity of this formula has been verified in many experiments [82]. Other authors have reached the same conclusion. Ostermayer, for instance, has shown that normal stresses at the anchor root surface increase to 2 to 10 times the stress attributable to the mass of the overburden [153]. It is nevertheless correct, as far as an approximate calculation of the required root length is concerned, to start with the average value for the shear strength at the root surface. On the basis of a steadily increasing number of experiments, it is possible to consider the following approximate values for xb: fine to medium-grained sand, moderately compacted to highly compacted (Fig. 13-18)

0.14-0.51

medium to coarse-grained sand and sandy gravel, moderately compacted

0.32 — 1.00

The value of zb increases markedly with soil compactedness, soil grain coarseness, and the coefficient of grain uniformity. The value of xb per unit length decreases with increasing root length, as occurs in hard rocks, and

129

therefore a disproportionately long root is uneconomical. L. Hobst obtained good results with the fixing of anchors in non-saturated coarse-grained gravel sands, using an anchoring length of only 3 m (see Fig. 9-12); he recommends, a root length of 4 to 6 m depending on anchor size and the quality of the soil. Ostermayer recommends an optimum root length of between 6 and 7 m. The graph in Fig. 13-19 shows ranges of experimentally obtained loadbearing capacities in relation to root length.

Fig. 13-18. Influence on friction at root surface of diameter and length of grouted root fixed in sand (fine to medium, strongly compacted to medium compact U = 1.6 — 3.1) [153]

bond-to-ground length [m]

2000 very dense

medium dense ^ dense dense medium dense medium dense

■i

Ύ

sandy gravel U-5-33

2

6

bond to ground length

fine to medium sand U- 1.6-3-1

medium to coarse sand (with gravel) diameter of grouted oody d010- 15 cm overburden ^^m

Fig. 13-19. Loading capacity of anchors in non-cohesive soils in relation to soil type and root length [155]; diameter of grouted body 10—15 cm; overburden height 4 m; U — coefficient of uniformity

130

In cohesive soils, the shear strength at the surface of long anchors is much lower than that of corresponding anchors in non-cohesive soils. The required anchor length must be determined on the basis of extraction tests, particularly in larger projects. However, the relationships obtained both in research and in practice up to now [154] may serve well as a basis for approximate calculations: a) Surface resistance to movement (friction) increases with increasing consistency and decreasing plasticity of the soil. The lowest computed values (0.05 to 0.08 MPa) of the average shear strength (friction) at the root/soil interface were found in stiff clays (I0 = 0.8 — 1.0) with medium to high plasticity; the highest values (around 0.4 MPa) were found in sandy silts of medium plasticity and very stiff to hard consistency (I0 = 1.25) (Fig. 13-20) [154, 53]. 600

very stiff to hard sandy silt with medium post-grouting plasticity very stiff to hard (marl) without } pst- groutina

^00 cloy £200

medium plasticity (marl)

very stiff without post-grouting stiff without post-grouting

6 * bond-to ground-length

c ^200

clay . medium to high plasticity

r

iS [mj

very stiff with post-grouting very stiff without post-grouting stiff without post-grouting t 6 bond- to -ground leng th [m]

10

12

Fig. 13-20. Surface friction occurring in cohesive soils for various lengths of grout fixing» with and without post-grouting [154]

Approximate surface resistance values may be also obtained for a given soil by means of pressiometric tests carried out in exploratory boreholes (see Chapter 9.2). b) The shear strength at the interface between test anchors and soil did not change with the fixing length up to a xh value of 0.1 MPa. The shear strength decreased slightly with fixing length at higher values of xb, but

131

this decrease may be neglected for the purpose of approximate calculations, the shear strength for any given soil being taken as constant. The maximum load-bearing capacity of an anchor thus increases, approximately in proportion to the root length. c) A similar situation to that in b) was observed for root diameters the range 9 to 16 cm. The ultimate surface friction value did not change within this range and the load-bearing capacity of the anchor increased proportionally with the root diameter. d) The surface friction of the anchor root can be increased substantially (up to 100 per cent.) by carrying out second, third etc., groutings, the effect being nearly proportional to the applied grouting pressure [102] (see also Fig. 13-20). It is important, however, that a rupture of the previous grout body occurs first, and that high pressure is maintained until the grouting is completed. In order continuously to improve the data available for deciding on anchor fixing length, intensive investigations in situ, in which anchors are tested to the point of failure in various types of ground, should continue. Data concerning the load-bearing capacities of anchors and the mean grout/rock or grout/soil bond strengths obtained in the course of research and in practice should be accompanied in all cases by exact information on the anchors, the anchoring method, and the characteristics of the ground. It will then be possible gradually to establish which are the critical parameters of ground conditions and how these relate to the load-bearing capacities of anchors. 13.2.1.3 Cohesion between grout and the steel components of anchor root It has already been stated that in rock anchorages the weakest part of the fixing is the bond between the grout and the anchor tendon, rather than that between the grout and the rock. The grout/steel bond involves three factors: — Adhesion, resulting from physical bonding between the surface of the steel and the adhering grout. Adhesion accounts for the first resistance that comes into play when both materials are stressed by shear forces. It disappears when movement of the root takes place. — Mechanical interlocking with the steel parts of the anchor owing to the presence of rolled-on ribs, threads, cavities and projections; the latter are moulded into the adhering grout body. Interlocking acts in combination with adhesion. — Friction, arising as a function of the clamping pressure and the roughness of the steel surface; the magnitude of the friction factor also depends on whether it is acting prior to movement along the contact surface,

132

in which case its value is higher, or whether it is acting during the course of movement, when the residual coefficient of friction at the surface is smaller. In the initial stages of stressing, adhesion and interlocking between uneven-surfaced elements are responsible for the integrity of the bond. When these factors are progressively overcome, beginning at the connecting point between the root and the tendon and moving towards the root end the friction factor comes into effect, in its lesser value as kinetic friction. The fixing strength obtainable by friction in this case is only a fraction ofthat provided by adhesion enhanced by interlocking; thus the friction factor in the case of long roots cannot be regarded as important in practice. Friction becomes a significant component of the fixing strength only where cable ends are fixed to conical anchor bases (see Chapter 13.3), and to some extent also where ropes and corrugated wires are used, as will be explained subsequently. The limit cohesion per unit surface area of the steel anchor is usually taken as being about one tenth of the compressive strength, or 4 MPa at most. The limit cohesion decreases with fixing length (Fig. 13-21), in the same way as grout/reck cohesion. This decrease arises from the non-uniform distribution of stress along the length of the fixing. On stressing the anchor, adhesion comes into effect initially at the proximal root end, while at the distal end it remains unexploited. When adhesion is overcome in the proximal section,

J

c

b)

Q)

a) kg/cm 150 -130 100 50 0

-

^^T"""

w

kg/cm Oospti) 50 _

>40

10 20 30 mm φ Tor anchoring length = 10 Φ

-

Φ 20 Tor

80 65

N^2

>37

10 20 30 W 50 cm I

Fig. 13-21. Relationship between anchoring length and cohesion between the shaped Tor bar and concrete (according to S. Soretz)

a process which is advanced by contraction of the anchor cross-section under tension, a slip takes place and most of the stress is gradually transferred to deeper sections, while at the proximal root end only the small effect of friction remains. It is clear that cohesive resistance does not act along the entire root length, except when the anchor is about to be torn out from the grout and only the effect of friction is resisting further movement. This behaviour of grouted anchors has been confirmed by L. Hobst and other authors in loading tests (Fig. 13-22), in which the stress distribution

133 4.0

SS· 7 · 0

&P

to 5>

1.0

^<^g4v£;

direction of- tension n distribution of 3-4 5-6 tensionmetenß . , , ,

0.0

0.5

7 i .

J_J

W

l_

1.5 depth of embedding in

72-75 2.0 concrete[mj

2.5

Fig. 13-22. Stress diagram obtaining under conditions of adhesion between concrete and steel, with different tensile forces in the latter

at the surface of steel bars and wires embedded in concrete was observed. When the extraction of steel anchor elements from concrete is carried out, care must be taken that the steel elements are embedded in concrete masses of sufficiently large diameter. In any case, embedding in cylindrical masses enclosed in steel containers must be excluded. With such an arrangement. the reaction transmitted through the support parts of the extraction equipment to a position near the zone of tension, or the reinforcing effect of the steel container, will affect the distribution of stress; this leads to values for the cohesion strength which are higher then any that could be considered in the design of tensioned roots fixed in rock. In comparing different anchor designs with respect to the necessary grouted fixing length, one can use the method recommended by the FIP-CEB regulations for prestressed reinforcing elements embedded in precast structures. According to these regulations, the length of an effective fixing is determined from the amount of contraction observed at the ends of the prestressed components embedded in the prefabricated structure, upon release from the prestressing equipment. The following relation is used: lk =

3.5Ea.Fa.A7lP,

where ΔΊ is the contraction distance of the reinforcing element into the precast structure after 7 days. The values obtained in this way for the fixing length are lower than those which could be recommended for anchor roots, the reason being that, when these measurements are made, the reinforcing elements are enveloped by fconcrete, and when the prestressing force is released, the cross-section of these elements expands. The expansion progresses outwards to the surface of the material as the cross-section reaches its original dimensions. The reinforcing elements have the effect of a plug which is pulled into a socket

134

formed by the adjacent concrete. On the other hand, the tensioning of anchor tendon is the very reverse process. The root length of an anchor, or more precisely, the length of the necessary wrapping of the steel anchor components with grout poured into the borehole, depends on the following factors: the anchor prestress, σ α , the surface area of the prestressed steel, the strength of the concrete, the rate at which prestressing is introduced into the anchor, the distance of the steel components from the root surface, the strength of the rock medium surrounding the root, the shape of the root (a conical root is preferable, because the shear stress vector at the surface of the steel can be increased by the vector of the product of compressive stress and coefficient of friction.), etc. Formulae for the calculation of the fixing length, in which the effects of the above factors are included, have been derived by various authors. These formulae, however, are capable of giving different values for the fixing length, the reason for these differences being that some of the coefficients included in the formulae are taken from the results of tests, the design and interpretation of which may differ considerably. When calculating the required root length of an anchor as a function of the fixing strength of the anchor steel parts to the grout, the equilibrium of forces acting at the limit state at the grout/anchor interface is considered: mP = n . π . dx . / . τα (see page 124). The formula also takes into account the anchor characteristics, particularly the tensile stress of the steel, σ'α, at the point of cohesion failure, and the sum total of the perimeters u of the anchor elements in contact with the grout: I = m —. K. u .τα K is a coefficient expressing the uneven distribution of shear stress along the fixing length corresponding to values of the ratio σ'α : u. It is determined experimentally. Mamontov, for instance, gives the value of K for steel ropes (Fig. 13-23) [125]. The summed perimeter, u, of the anchor elements is determined for an «-fold bundle of wires or bars as the «-multiple of the perimeter of an 1.8 1.6 lit 1.2 1.0 0

Ί 1L

Pt

/<9

MI cm3

22

2f

28

Fig. 13-23. Coefficient, K, expressing the uneven distribution of shear stress, τ, along the anchor fixing length

135

individual element (u = nn . dt). For steel ropes, u is taken as the total outer perimeter of the external layer of wires, which for a 7-strand rope is approximately four times the perimeter of an individual strand of diameter d\ (u7 = An . d')\ for a couple of 7-strand ropes, ulml = In . d\ etc. The cohesive strength, τ α , for concrete structures is taken, as stated previously, as one tenth of the compressive strength of the grout, up to a maximum of 4 MPa. Considering the difficulty of determining safely the quality of concrete made by grouting the anchor borehole (in which standard conditions for the preparation of good quality concrete cannot be established), it is safer to rely only on the lower values within the range of τα (1—2 MPa). It should be noted that the strength of the grout affects, to a relatively small extent, the required fixing length of plain bars and wires in concrete, although, theoretically, the fixing length should be inversely proportional to the grout strength. Grout strength develops at a considerably slower rate than the strength of concrete, and also more slowly than is assumed in some Standards. In tests on concrete structures it appears that the cohesion strength can still grow at compressive strengths of the concrete up to 30 or 40 MPa, while at higher compressive strengths the growth is negligible [120]. The strength of concrete, however, is fully exploited where anchors are composed of deformed bars with which the concrete interlocks. In the latter case cohesion becomes the most important factor, acting permanently and not becoming lost by creep of the concrete. The root length necessary for an anchor increases with the number of anchor elements. This increase is not, however, proportional to the number of elements, as will be explained further on. For practical calculations of the fixing lengths of bars, values for the admissible stresses, τ α , acting on the cohesive bond (as recommended in individual national Standards) are substituted in the following formula: I, = m

P

where ur is the sum of the perimeters of the individual bars multiplied by kr, a coefficient, expressing the effect of the number of bars on the magnitude of /. The British regulations, for instance, give the following values for this coefficient: for two bars fixed in a common channel (borehole), kra = 0.8; for three bars, krt3 = 0.6; for 4 bars, krA = 0.4. Where anchors are composed of larger bundles of bars, the ends must be provided with a base. The fixing length of wires in the grout of the anchor root is usually calculated in terms of a multiple of the cross-section of a single wire. For single wires the recommended fixing length is from 100 to 200 times the diameter. The need to increase the fixing length, as the anchor usually contains a larger number of wires, is at least partially eliminated in the majority of

136

wire anchors by corrugating the wires in the root section, usually by pulling them through a system of clamping and spreader rings, or by crimping them in the workshop (see Figs 12-7, 13-29). When corrugated wires are pulled, frictional forces come into effect, as shown by a simple diagram of the forces acting on a stressed wire (Fig. 13-24).

Fig. 13-25. a) — Stress distribution in area of contact of rope, b) — Dependence of the friction component, p, on the angle a (see text) 1 ~ a = 10°, 2 — -a = 12°, 3 — a = 12° (limit state), 4 — a = 16°

The necessary fixing length of strands in the grout of the anchor root can be taken as lying within the range of 30 to 50 times the cross-section. This relatively low value arises from the favourable nature of the rope surface. The concrete forming in the depressions among the helically twisted wires is stressed by a friction effect arising from the normal component, N9 of the force P. N increases with the angle, a, between the axis of the outer rope strands and the axis of the rope itself (N = P. sin a). As an example, if a increases from 12° to 16°, the area of contact and the initial resistance to displacement increase by 1.75 %, but the frictional effect, as a component

137

of cohesion coming into play after adhesion failure, grows to 150 p.c. (Fig. 13-25). It appears from the above that for interlocked strands of the Dyform type, the fixing length must be increased by 20 to 30 %. The fixing strength of an anchor composed of a bundle of ropes decreases in the same way as that of anchors made of bundles of bars decreases in comparison with single bar anchors, with the difference that the coefficient of decrease of efficiency has a higher value (e.g. for 2 ropes, kr>2 = 0.85). This means that the adverse effect of increasing the number of anchor elements in extending the necessary fixing length is smaller for ropes than it is for bars; moreover, this effect can be further reduced by corrugating the fixed sections of the ropes in the same way as wire anchors (Fig. 13-26). Ropes with a cross-section of more than 20 mm and comprising a larger number of strands should be spliced in the root section to ensure thorough contact with the grout (Fig. 13-27). In this case the method of fixing is the same as that used for a bundle of corrugated wires (see Fig. 12-7).

Fig. 13-26. Arrangement of steel strands in the fixing section of VSL rock anchor

Fig. 13-27. Prising apart the end of a Hercules rope of nominal load-bearing capacity 7.5 MN (Reconstruction of Bystficka Dam, Czechoslovakia)

138

When the root length of any type of anchoring is being considered, it should be borne in mind that in the course of time the fixing will become loosened by rheological phenomena, particularly where anchors are prestressed shortly after grouting. For this reason it is recommended that the safety factor for failure of cohesion between the grout and the surface of the anchor steel components should be greater than 2, and that the calculated length be increased by 50 to 100 per cent. A basic condition for reliable fixing is also a high strength, or noncompressibility, of the ground in the vicinity of the root. This is aided by grouting under pressure. 13.2.2 Technology of fixing anchors by grouting The fixing of anchors by grouting is the most highly developed fixing technique at the present time. The aim of every grouting system is to drill the borehole as quickly as possible, insert the assembled anchor tendon into the borehole with ease, and then perfectly grout the borehole and/or the immediate rock or soil so as to create a load-bearing root and reliable anticorrosive protection of the tendon throughout the service life of the anchor. The appropriate procedure is selected according to the type of ground involved, and the design of the anchor. Generally it has to be borne in mind that the weaker the rock or soil and the smaller the assumed cohesion between the ground and the anchor root, the more exacting are the requirements placed on the anchoring technology if reliability and economy of installation are to be ensured. 13.2.2.1 Short bar anchors (bolts) Short steel bars, usually shaped (rebars) and with threaded external ends, are fitted into prepared boreholes in rock which are then filled partly or completely with grout or mortar. If they are fixed only at the remote end of the borehole, they can also be prestressed. If, however, the entire length is embedded in grout, the bolt remains unstressed and reinforces only the rock mass in the vicinity of the excavation. Compared with mechanically fixed bolts, those embedded in grout are very much cheaper and can be used with success in softer rock types; however, the fixing of grouted bolts into boreholes is more complicated and it takes longer for these bolts to be brought into use, since the strength develops with hardening of the grout or mortar. In case that an accelerator is added the mortar may start to harden after a few minutes and a sufficient strength is reached in two or three hours. Anchoring bolts of rebars, fixed throughout the borehole length by grout, which is injected through a special tube inserted down to the borehole

139

bottom and pulled out step-by-step with the progress of grouting, are called SN-bolts in Europe after Store —Norfors, a place in Sweden, where they were used for the first time. Bolts, enveloped by grout along their entire length, are more effectively protected against corrosion. Correct positioning in the centre of the borehole must be established by means of suitable spacers attached to the bolt. The bottom part of the borehole must be often de-aerated to ensure its perfect filling by grout. This is provided by a plastic tube of small diameter, inserted together with the anchor bar down to the bottom of the borehole, or by the bar itself, which is hollow in this case. The so-called dry bolts without prestress are short rebars fixed in soft rock throughout their full length only by the grip of the rock, without cement. They are driven mechanically into the borehole whose diameter is smaller by 2 to 3 mm than that of the bar. This type of anchorage was successful, TABLE 13-111 Bolts fixed with cement mortar into hard rock Admissible Loading loading capacity form = 1.5 [kN] [kN]

Bar diameter1) [mm]

10 20 30 40 50 60 70 80 90 100 120 140 160 180 200

8 12 14 16 18 20 22 25 25 28 28 32 36 36 36

1

15 30 45 60 75 90 105 120 135 150 180 210 240 270 300

Threading on bar

M8 M12 M 14 M16 M18 M20 M22 M24 M24 M27 M27 M30 M33 M33 M36

Fixing length2) [cm]

Borehole diameter2) [mm]

24 32 41 48 53 58 61 61 69 69 82 86 85 96 106

11 17 20 23 25 29 32 36 36 40 40 45 52 52 52

Grout volume*) [dm3]

0.02 0.05

0.1 0.15

0.2 0.25 0.35 0.45

0.5 0.6 0.7 0.9 0.8 1.4 1.6

) For deformed (shaped) steel bars (rebars) of yield point 400 MPa. For the analysis, the diameter of the shaft of the bar (excluding the thread) is considered. Taking the strength of the bar/concrete bond as 2.5 MPa. The figures represent the diameter of the drill bit. The necessary diameter is determined 2 ) from the required safe load-bearing capacity of the fixing between the anchor and the 3 ) rock, taking an average cohesion of 1.8 MPa between the concrete and the rock of the borehole 4 ) The figures represent 4/3 of the theoretical free space of the borehole

140

for example, in the anchoring of the faces of excavations in clayey shales for the Prague Underground railway, or in the excavation of a gallery in much saturated shales under the Rhine bed in the GFR. Some important data pertaining to bolts grouted in strong rocks, with the fixing length necessary for taking the admissible tensile force, are listed in Table 13-111. The data are compiled on the principle of full exploitation of the materials used, taking the safety factor a s m = 1.5. For the current fixing of short prestressed bolts, 300 to 1,000 cm 3 of grout must be injected into the borehole. To place these limited quantities of grout in upward-directed boreholes, simple hand-operated aids as discussed in Section 15.3 are used. 13.2.2.2 Long anchors in hard rock Long anchors are normally fitted into boreholes with solid walls. Where layers of unconsolidated sediments overlie the bedrock, the boreholes must be cased. It is sometime expedient to employ the Duplex system, where drilling is effected by the central rods as well as by the external casing pipe (see Section 14.2). Prior to insertion of the anchor into the borehole, the section of tendon destined to transfer the tensile forces to the ground is arranged according to the system used, the anticorrosive protection of the tendon is checked, and a pipe and sealing element for single or repeated grouting are attached to the anchor. Unless increased anticorrosive protection of the anchor is stipulated, the steel parts of the tendon in the fixing section are left without any protection. The prestressed sections of the tendons are simply covered with a plain flexible tube or an insulating coating to protect them from contact with the grout, as well as from possible aggressivity in the ground medium. The anchor is grouted by means of a single grouting pipe from the deepest point upwards, applying low pressure until the grout completely fills the borehole (Fig. 13-28). In densely jointed or soft rocks the grout must be pumped in under pressure. To achieve this, the fixing section of the borehole is closed off by either an elastic seal (Fig. 13-29), or a linen sealing bag fixed on the anchor tendon (sec Fig. 13-43). Under particularly exacting conditions such as those encountered in soils, a grouting sleeve pipe is employed (see Fig. 13-40). In such cases the grouting is repeated, either once, or possibly several times. The arrangement of the anchor tendon in the fixing section varies according to the type of steel elements employed. Plain wires and strands are usually undulated and spread alternately by means of clamps and spreader rings (see Fig. 12-7), sometimes also by cold shaping of the wires (see Fig. 13-29). The splicing of steel ropes together with thorough cleaning of the wires is often

hex nut anchor plate

*ent and grout tube

sheathing

Fig. 13-28. Details of temporary rock anchors A — Dywidag single bar anchor, B — VSL 9-strand (7 wires in each strand) anchor

qement gnoat

Section A-A

smooth sheath

grout tube*

smooth sheath

threadban

142

sufficient (see Fig. 13-27). For shaped bars (as used, for example, in the Dywidag and Bauer systems), the steel/grout bond suffices without any further measures. The centering of the bar in the borehole must be established with the aid of spacers which are attached to the bar and shaped so as to glide easily against the borehole walls when the anchor is inserted. The spacers must ensure that there is a minimum grout layer of 2 cm covering the steel.

Fig. 13-29. Polensky & Zöllner anchor system (PZ). The fixing section of the borehole is sealed off by two leather collars a) — anchor root, b) — tendon, c) —anchor head; 1 —cable, 2 — plastic pipe or insulative wrapping, 3 — sealing collars, 4 — grouting pipe leading to fixing section of borehole, 5 — grouting pipe for tendon section, 6 — de-aeration pipe, 7 — helical reinforcement

The fixing efficiency of anchors embedded in hard rock or soil, provided that the adhesive properties of the grout have been fully exploited, can be increased by ensuring that the anchor tension acts at the distal end of the root rather than the front (Fig. 13-30b). This is achieved by forming the long insulating coat not only along the part of the tendon outside the root, but also along the entire length within the root. In this way, adhesion between the surface of the wires, bars, or strands of the anchor and the grout is disposed with, but adhesion between the grout and the borehole wall remains unimpaired. The missing component of the fixing system (adhesion between grout and steel) is substituted by an anchor base formed by a distribution plate or a head piece abutting against the filling of the borehole. When the

143

Fig. 13-30. Basic types of anchor fixing involving a concrete root, and the stress occurring in the root a) — gradual stress development caused by tension, b) — stress caused by concentric compression at base, c) — shear and compression stresses distributed over a larger area

tendon extends to this base, compression of the hardened grout of the anchor root takes place on prestressing of the anchor. The compression induces a transverse stress in the grout above the root base, tending to cause radial deformation, that is, giving rise to an expansion pressure against the borehole wall. The transverse stress increases with the force exerted by the base in the loading direction of the anchor; this occurs according to the Poisson's ratio v, of the grout:
G»

If for a compact concrete the value of the Poisson's ratio is assumed to be 0.17 the transverse stress created by the force of the anchor base has a value of 1/5σΑ. Let us consider by way of example, an anchor of 1,000 kN tension embedded in a borehole of 15 cm diameter, with a base of the same diameter, and a 6 cm diameter tendon. In the concrete immediately above the base of this anchor, a stress of 68 MPa is developed with a transverse pressure on the borehole wall of 13.5 MPa. This transverse pressure increases the fixing strength of the anchor in any type of rock; therefore, it is not advantageous to contain the radial pressures acting in a compressed root within any form of reinforcement, such as a helical coil or a coaxial steel pipe similarly fixed to the anchor base, as some authors recommend for soils (Fig. 13-31). A pipe transfers the pressure of the base over a greater length of the root, and this reduces the stress in the concrete immediately above the base; at the same time, however, the transveise pressures are contained by the pipe and the advantage of expansion of the concrete so that it is forcibly pressed against the borehole wall is lost. The concentration of stress immediately above the anchor base creates the possibility of the root concrete being crushed; however, this outcome

144 ΙΛ3

0.8 r-G-7

if a* ^0.3 $0.2

0.7 0

\YaS \s*

/

b.

^

^Ί

type of the ba

.^ ff= Ί

Fig. 13-31. Relationship between the ultimate loading capacity of an anchor in soil and the length of a steel cylinder fitted at the base, according to H. Bendel [15] a — sand and gravel, b — fine sand

*/"Ρ

1 2 3 * + length of the base [mj

cannot, apart from exceptional cases, reduce the fixing strength of the anchor as a whole. Indeed, the Poisson's ratio v, for crushed concrete is less than 0.17 and takes a value similar to that for condensed sand or rock fragments v = 0.25—0.33, so that the transverse pressure of the grout increases to a value within the range ah = 0.33συ to 0.5συ. In the above example of an anchor of 1,000 kN load, ah would then be 22 to 34 MPa. The radial pressure exerted on the surrounding rock by both the anchor root and the borehole filling above the root, decreases progressively towards the borehole mouth. In spite of this, the overall effect of this pressure is to produce the chief component of the resistance to extraction of the anchor from the ground, the most important characteristic of which, particularly with respect to cohesive and loose soils, is the angle of internal friction. Thus L. Hobst recommends that for anchoring in soil, the space of the borehole around the anchor above the base should be filled with a wall compacted gravelly sand showing a continuous granulometric curve. This is preferable to concrete, unless, of course, concrete is essential for the basic anticorrosive protection of the anchor. The outward pressure of the borehole filling as a result of its being compressed by the tensioned anchor base, makes it possible to exploit fully the internal friction of the soil and gain a high fixing strength. The validity of this analysis has been demonstrated in laboratory and field tests. Where anchors are installed in strong rocks it is technically preferable to fill the borehole space around the anchor (i.e. above the base and along the length of the tendon) with concrete, because the concrete, on account of its good adhesion to the borehole walls, facilitates full exploitation of the rock shear strength, which is the dominant component of the fixing strength in strong rock; it also protects the anchor against corrosion (see Chapter 18). In tensioned roots which are fixed to the tendon along the entire fixing length of the root (see Fig. 13-30a), there appear transverse cracks close to the root front, and these can be the cause of steel corrosion. For this reason the roots of permanent anchors designed in this way are protected by a special plastic tube extending into the fixing section (Figs. 13-32, 13-33); the tendon section inside this tube is carefully grouted or filled with resin, and then sealed

smooth sheathing

Section A-A

corrugated sheathing

Fig. 13-32. Arrangement of permanent rock anchors with double anticorrosive protection A __ Dywidag single bar anchor, B — VSL 9-strand anchor (see Chapter 18)

///

cap anticorrosive compound hex nut

vent andgnout tube

anchor plate

plastic coated and greased strands

146

either before or after insertion of the anchor into the borehole. A strong connection of the tendon thus protected with the grout of the root is ensured by the undulatory surface of the plastic tube. Even if transverse cracks occur in the filling of the tube, usually after the tendon has been loaded, they are no longer dangerous since the> are covered by the undamaged wall of the plastic (Polyvinylchloride or polyethylene) tube (see Chapter 18). 13.2.2.3 Grouted anchors in loose soils Many procedures have been developed for obtaining anchorage in loose soils by grouting; an anchor root of the required load-bearing capacity can be formed without difficulty in these soils, by grouting. BBRV

'

CON A ~ Sol

Fig. 13-33.-I. Cross-section of tendon and fixing sections of BBRV wire anchors and CONA-Sol strand anchors for temporary and permanent anchoring (documentation Stahl Ton, Switzerland) A — temporary anchor, B — permanent anchor, C — monitored (permanent) anchor, 1 — borehole, 2 — wire dia 7 mm, 3 — strand dia 12.7 mm (0.5 in.), 4 — protective plastic tube (smooth), 5 — pipe for primary grouting, 6 — pipe for additional grouting, 7 — grout outside the protective tube, 8 — grout inside the protective tube, 9 — permanent plasticity compound, 10 — spacer, / / — corrugated plastic protective tube

147

The procedure develop« d by the Swedish company Hagconsult AB (Stockholm) is the simplest. The borehole need not be cased and the same drillrod is used for boring, g outing and anchoring (Fig. 13-34). The drillrod set which is made of highgr de steel (32 by 16 mm outside and inside diameters, tensile strength 520 kN), is taken to the required depth by means of a light wagon drill. In the final drilling stage, pressurised grout is substituted for the drilling fluid, and the end of the drillrod, together with the BBR V

"

CONA - Sol

Fig. 13-33.-II. a — double anticorrosive protection by means of grout and plastic tube, b — protection by means of 20 mm grout layer, c — extra protection by means of steel tube

148

V

Ot

■ O 1

O / O

c? o /

>

4? o -

O o

■ o

<6

0

'

:

' ffl< <*

Fig. 13-34. Mounting and injection of an anchor composed of drillrods (Hagconsult system)

V

SJ

,>

Ϋ* ■ /

/ ' Λ

/

/

bit, is grouted in the soil thus forming the anchor root. The outer end of the drillrod set is then fitted with a threaded endpiece to take the tensioning nut, and after the grout has hardened the anchor is prestressed in the usual way. However, such an anchor can only be a temporary one, as the tendon is not protected from corrosion. Uninterrupted grouting of a perforated anchoring pipe directly rammed into the soil is the basis of the MV system (Fig. 13-35). It is most frequently used in saturated soils of high ground water level for the transfer of tensile, but more often compressive, forces in emergency work, and in making improvements to the load-bearing capacity of foundations. Such loadbearing elements of small diameter are often referred to as micropiles. In other anchoring systems for use in loose soils, the anchor tendons are placed in cased boreholes, drilled either by rotation or percussion. Very often the casing pipes are vibrated or rammed into a loose soil, in which case the end of the pipe is equipped with a shoe which remains in the ground. The entire anchoring procedure is shown in Fig. 13-36, illustrating the type patented by the West German Bauer Company. This system uses tendons

149

made of shaped bars which are provided with anticorrosive protection and placed in cased boreholes. The root and the free tendon length are grouted from above through the casing pipe, which is gradually pulled out of the

Fig. 13-35. MV anchoring system 1 — rammed anchor pipe, 2 — lost shoe through which grout is squeezed, 3 — grouted soil, 4 — grout input, 5 — ramming direction, 6 — water level in stressed ground, 7 — completed anchor

Fig. 13-36. The steps of establishing an anchorage in soil by the method of the K. Bauer Co. (GRF) 1 — a cased borehole of 70—150 mm diameter is drilled with a shoe at the inner end, 2 — the tendon with protective tube is inserted and connected with the shoe, 3 — the casing pipe is extracted from the borehole with simultaneous grouting of the fixing section, 4 — a tensile test of the anchor is carried out 6—8 days after grouting, 5 — the anchoring head is connected to the anchored structure, and the anchor is prestressed with the required force

150 ground at the same time. If the tendon, after insertion into the casing pipe, is screwed into the deposited shoe and provided with a protective coating along its entire length, then the root is favourably stressed by compression from below when the anchor is loaded. Fig. 13-37. Weber anchoring system used by Stump Bohr AG 1 — assumed shear surface, 2 — protective outer wrapping of the anchorage, 3 — double wrapping with perforations and steel grouting base (4), 5 — prestressed bar (tendon), 6 — grouted zone in a permeable soil, 7 — stress diagram indicating the varying magnitude of the shear stress at the surface of contact between the anchor base and the concrete of the anchor root

The Swiss company Stump Bohr AG of Zurich [15] uses the Weber system, in which the force from the tendon is directly transferred to the root by compression. A protective tube, made of thin corrugated sheet or plastic provided with a long grouting steel base at its-lower end, is inserted into the cased borehole. The steel base is of double thickness, and the external surface is perforated in the section designated for grouting (Fig. 13-37). A grout pipe is then screwed into the grouting base. As the casing is extracted from the borehole, the soil is grouted in the fixing section, and an anchor root is formed at the lower end of the borehole in conjunction with the grouting base. When grouting is completed, the grout pipe is unscrewed from the base, and a bar representing the anchor tendon is screwed on and prestressed. The bar is free between the lower and upper threads, and is stressed only by tension. In a temporarily anchored structure, the tendon may be taken out when it is no longer required; the outer protective sheathing and the base are not recoverable. In the case of a permanent anchorage, the prestressed bar is grouted in the sheathing or wrapped in advance with a suitable material to protect it against corrosion. A more recent anchoring method used by this company is that shown in Fig. 13-38. A bar anchor provided with an insulative wrapping and a steel compression member at the base is inserted into the borehole and is grouted through the casing pipe, as in the Bauer system. Some anchoring systems used m so:ls successfully adopt the technology

151

of rock anchors, with small modifications (Fig. 13-39). These anchors obviously are placed into cased boreholes and are better protected against corrosion.

Fig. 13-38. Recently introduced anchoring system used by Stump Bohr AG 1 — the cased borehole is prepared, 2 — the borehole is filled with grout from the bottom upwards with the aid of a pipe, 3 — the assembled anchor is inserted, 4 — the anchor root is grouted with simultaneous extraction of the casing pipe from the borehole, 5 — the anchor is tested and prestressed after hardening of the grout

Fig. 13-39. Main types of Dywidag soil anchors A — single bar anchor with simple anticorrosive protection, including*post-grouting system, B — single bar anchor with simple anticorrosive protection, recoverable, C — multiple bar anchor (3 to 9 threaded bars) with simple anticorrosive protection, D — single bar anchor with double anticorrosive protection

154

A significant contribution to the technology of long grouted anchors in soils is represented by the procedure introduced by the French Soletanche Company (IRP system). By using a sealing bag to separate the anchor root from the free section of the anchor tendon and a collared grouting tube (see Fig. 13-40), high-pressure grouting of the root, and also of further sections of the anchor is made possible and repeated grouting can be carried out whenever needed. The sealing bag replaces the rubber or leather seal located in the borehole when grouting in rock. The sealing jute bag is usually 2 m long; it is slipped on to the anchor at the upper end of the fixing section and tightly fastened to the anchor at both ends. When the anchor has been inserted into the cased borehole and the casing pipe pulled out, the bag, filled with grout under low pressure (up to 0.5 MPa), presses against the borehole walls and isolates the fixing section (Fig. 13-40). The collared tube is a large diameter PVC tube provided with lateral apertures at 1 m intervals, these being covered on the outside by rubber collars (Fig. 13-41) so that the grout can flow from the tube into the borehole, but not in the reverse direction. The collared tube is attached to the anchor tendon and passes through the sealing bag (see Fig. 13-40). A steel grouting pipe of smaller diameter is passed into the collared tube; the perforated section of this steel pipe, delimited by two seals, is located at the opening of the collared tube at the necessary depth, and the borehole is then grouted under pressure (see Fig. 13-40). When grouting at this position is completed, the grouting pipe and the collared tube are flushed with water to keep them clean, and grouting can then be carried out at the next aperture. When an anchor is grouted by means of a collared tube in this way, the sealing bag is filled first. When the grout in the bag has set (after 6 to 12 hours), grout is forced into the fixing section by stages, starting from the bottom of the borehole; where the anchor has been prestressed, the free tendon section is grouted also. If the required load-bearing capacity of the root has not been attained, grouting of the root is repeated using a higher pressure which ruptures the former hardened grout and forms a better and stronger fixing with the surrounding soil (Fig. 13-42). The collared tube principle is used by many companies for anchoring in soils. The collared tube is placed inside the tendon among the wires, strands, or ropes, or it is fastened at the side in the case of a bar anchor (Fig. 13-43); alternatively, the anchor tendon is inserted into a sufficiently large collared tube after all the stages of the grouting o r t h e soil have been completed. This latter arrangement does not allow for subsequent re-grouting of the soil, but it furnishes additional protection against corrosion of the tendon. The Soletanche Company offers a bar anchor which is designed on this principle, and is perfectly protected against corrosion; it has a compressed grouted

155 i-1 ni.

o

r

'o

Fig. 13-40. IRP anchoring system with sealing bag and collared tube developed by the Soletanche Company

Fig. 13-41. Collared grouting tube in fixing section (holes covered with rubber sleeves) a) — with an internal overpressure the sleeve allows the grout to escape into the surroundings, b) — with an external overpressure the sleeve seals the holes in the tube

156 Fig. 13-42. Cross-section of a regrouted body showing the 7 steel bars of the tendon and the centrally located collared tube for grouting. The white lines are cracks filled with post-grouting material (documentation of K. Bauer Co.)

Fig. 13-43. Long bar anchors of Vodni stavby Czechoslovakia fitted with polyethylene collared tubes, centering metal sleeves, and linen sealing bags

root in the soil (Fig. 13-44). The Tubfix system employs a steel collared tube of sufficient strength to act as the anchor tendon itself. It has only a simple anticorrosive protection by a grout layer which cannot be increased by applying a protecting coat, because the fixing strength would then be reduced. 13.2.2.4

Grouted anchors in cohesive soils

The fixing of anchors with long roots in cohesive clayey soils is more difficult to achieve compared with anchorages in loose soils. It is successful only when the right technology is applied and when the following principles, verified both by theory and in practice, are observed: 1. The boreholes should be larger than 10 cm diameter. The load-bearing capacities of anchors in cohesive soils increases proportionally with the cross-sectional area of the root, as research results indicate (see Section 13.2.1.2).

157

Fig. 13-44. Anchor installed inside a collared tube which provides for grouting and anticorrosive protection of the bar (Soletanche system) 1 — external bar end, 2 — nut, 3 — retaining wall, 4 — smooth protective tube in the free length section, 5 — coupling, 6 — grouting rubber collar, 7 — collared grouting and protecting tube, 8 — anchor bar tendon, 9 — bar fixing base, 10 — cap, 11 — protecting layer of grout, 12 — grouting hole, 13 — strengthened tube in the fixing section

2. The boreholes should by drilled with rotary drills, and cased simultaneously to prevent loosening of the borehole walls and excessive exposure of the clayey soil to the flushing water. 3. The fixing section should be thoroughly cleaned with compressed air and the root should be grouted as soon as possible after completion of the boring operation (within 12 hours). It is essential to place centering spacers in the fixing section of the anchor tendon. 4. The grout should be as thick as possible (maximum w/c, 0.4), and the pressure as high as possible without rupturing the soil. The pressure should be maintained until grouting is completed. 5. There should be a facility for re-grouting, allowing the load-bearing capacity of an anchor in a cohesive soil to be increased by up to 100 %. 6. Anchors in which the tensile stress is transmitted right up to the end of the fixing section (compressed root anchors) should be used. By applying the correct procedures good results can be obtained, even in cohesive soils. For example, ultimate loads of more than 500 kN were obtained using anchors with long tensioned roots fixed in Frankfurt clays. The permanent deformation of the clays was up to 20 mm, the consistency was 0.8 to 0.9, the peak strength parameters were φ = 20°, c = 0.02 MPa, and the clays were very sensitive to water [23, 24].

158

13.2.3

Fixing of anchors with synthetic resins

The cohesion developed between synthetic resins and strong rocks is two to three times greater than that between grout and rock. Another advantage of resins is a quick setting time which can be selected from within a range of several minutes up to several hours. Resins also exhibit excellent resistance to the corrosive effects of the rock medium and the dynamic effects of shocks. The disadvantages are the high cost compared with grout, and some degree of dependence of the setting time on the temperature of the surrounding rock. The placing of anchors in resin and the filling of boreholes are also somewhat more exacting. For these reasons, resin has only been used up to now for fixing short bar anchors (bolts) in situations where the face of a rock excavation has had to be reinforced quickly and effectively. Only exceptionally has resin been used for fixing long anchor roots in rock, since rapid setting is not essential, and the profit gained from any reduction in the setting time before commencement of the prestressing does not cover the increase in cost involved in using the resin. The best resin characteristics for the fixing of anchors have been found to be those of non-saturated polyester resins. These are the least sensitive to low temperatures and moisture coming from the medium with which they are in contact during setting; they are also capable of taking a large quantity of inorganic filler. They are thixotropic, which reduces their viscosity in the course of insertion of the bolt and minimizes the tendency of uncured resin to run out of the borehole during the installation work. The curing takes place also under water after mixing of the resin with the appropriate catalyst on the basis of peroxide. The quantity of catalyst used governs the curing and setting times, the latter being about five times longer than the former. The curing time of the currently used mixtures ranges from 1 to 30 minutes at a temperature of about 20 °C. Lower temperatures significantly retard curing and setting, while higher temperature accelerate them. The quantity of resin used, the time taken to prepare the mixture, and the ambient temperature, together determine the viscosity of the mixture. The longer the borehole, the more difficult it is to place the bolt, and the lower the temperature, the lower must be the viscosity of the resin so as to achieve thorough mixing with the catalyst; this is very important for proper setting. A perfectly set polyester resin was found to have the following values for the limit cohesion, Tmax, with the main rock types (according to tests conducted at Imperial College, London):—Table 13-IV. In designing a bolt fixed by resin, it can be assumed that the maximum load-bearing capacity increases uniformly with the length of the fixing in the borehole. The attainment of the full strength of the resin as a function of the curing time at 24 °C, for mixtures with different initial setting times, is shown

159

in the graph in Fig. 13-45; the relationship between curing time and temperature is shown in Fig. 13-46. Both graphs are published by the American Du Pont Company, Bolts fixed with resin may be prestressefr so as to exert a pressure on the rock, non-prestressed, in which case the rock is reinforced by the dowel effect, or they may fulfil both of these functions (Fig. 13-47). To be able TABLE 13-IV Limit cohesion of polyester resin with rock Rock type

Limit cohesion

Rock of average compressive strength 5 MPa (claystones, siltstones)

1.2—1.6 MPa

Rock of average compressive strength 14 MPa (coal, shales, marlstones, sandstones)

1.6—3.0 MPa

Rocks of average compressive strength 50 MPa (sandstones, limestones)

3.0—5.0 MPa

Rocks of average compressive strength 100 MPa (igneous rocks such as granite)

4.0—7.0 MPa

36 \l5-30 min.resin

32 28 2k

JO

100

*16

-15-30 min. resin

£60

1 20

12

^5-1 0 min. nes/n

\i \J

f

I

1-2 min. resin

10 20 30 40 50 60 cure time [min]

Fig. 13-45. Relationship between final strength and curing time at 24 °C for Fasloc polyester resins of different setting times

510 min. resin

8 \l-2n im. resin

70

<

0 35 H5 55 65 85 95 °F <65 7.2 12.7 18.2 23.7 29.1 3ϊ7°0 temperature

Fig. 13-46. Relationship between curing time and temperature for Fasloc resins of Du pont Co.

160

ll»uu*-»___\|

KUttwttwwaawwwwwua

Fig. 13-47. Three modes of application of resin-anchored rock bolts, using fast (dark), and slow (stippled) setting

iensioned bolts with resin point anchorage

tens zoned bolts fully resin anchored and grouted

mmmmm unten sioned dowels fully resin grouted

to prestress the bolt it must be fixed only at the inner end of the borehole. In this case the bolt is not protected against corrosion. However it is possible to install bolts using fast setting resin for the fixing of the bolt at the inner end, and slow setting resin as an anticorrosive protection for the remainder. Such a bolt is prestressed after the curing of the fast setting resin, but before curing of the slow setting resin. Non-prestressed dowels are fully embedded in resin and are loaded only when movement of the rock takes place. Formerly the resin used to be transferred into the borehole by forcing the compound mixed, so as to have a longer curing time, with a power pump or hand pump (see Fig. 15-6). Nowadays there is a widespread use of cartridges "^nich are filled in the factory and which contain resin and catalyst in, separate plastic wrappings ready for use on the spot (Fig. 13-48). S u 4 cartridges can be stored for more than six months. They are usually 30 cm long with a diameter of 20 to 40 mm for boreholes of 22 to 50 mm diameter. They are supplied, for example, by Celtite (Seifix), Du Pont, Williams, Titan, Dywidag, Nobel Cyanamid Meynadier, Torque Tension, Lenpir et Mernier, and many other manufacturers. The entire procedure for fixing a bolt in resin" and prestressing it is summarized in the following paragraphs. The procedure should be followed very carefully to ensure success with this method (Fig. 13-49). Fig. 13-48. A resin cartridge for insertion into an anchor borehole

161 Fig. 13-49. Installation sequence of resin-fixed bolt 1 — insertion of resin cartridges, 2 — insertion and spinning of bolt, 3 — stressing of the bolt after the resin has cured

V 1. The drilled borehole must be carefully cleaned. It should have the smallest diameter compatible with the selected bolt and cartridge size. 2. The appropriate resin cartridges are inserted into the borehole. Fastsetting cartridges are placed at the inner end of the borehole so that tensioning of the bolt is not delayed, while slower setting cartridges are placed in the remaining length of the borehole to complete the protection of the bolt. 3. The number of cartridges needed, given the diameters of the borehole and bolt and the length of the latter, are indicated by each manufacturer in the instructions provided. Damaged or partly set cartridges should not be used. 4. The bolt is inserted by hand and then spun with a drilling tool so that it breaks through the cartridges in the borehole. The spinning should continue for 30 to 60 seconds after the bolt has reached the bottom of the hole to ensure thorough mixing of the ingredients of the cartridges. The bolts should be shaped so as to exploit the full strength of the resin. 5. Rotation of the bolt is stopped, and it is then pushed inwards with the maximum thrust available from the drilling tool, and held in this way for several minutes until the fast setting resin sets. 6. The bearing plate is mounted and secured with a nut. Wedge washers are used if the plate is not resting at right angles to the bolt. 7. The bolt is tightened with the nut and stressed with a hydraulic jack, torque wrench, or impact tool (see Chapter 17). This is done after the fast cartridges have been allowed to set (5 minutes), but before the slow cartridges have set (20 minutes). One hour after installation, the slow resin has set thus locking the tension of the bolt and giving complete anticorrosive protection together with permanent reinforcement of the rock. Bolt anchors installed in this way maintain their tension in spite of vibration, or blasting which may be carried out nearby. When such bolts are

162

loaded beyond the failure point of the bond, rupture of the anchorage is not sudden, but rather the bolt begins to yield very slowly; in fact, the resistance to extraction may increase as the plug of material, which is moulded to the shape of the borehole, starts to move and thus pick up fine particles from the wall of the borehole. Fiber glass anchoring rods Reports have appeared during the last decade on the successful employment of glass-reinforced plastic bars as prestressed reinforcement structures, and therefore also as the elements of anchors. The advantages of such bars are their resistance to corrosion and the ease of fixing, not only in rock but particularly also with the use of resins at the anchor head. (Fixing with grout is not very efficient and is possible only when the bars are threaded (Fig. 13-50) or a base piece is attached at the end).

Fig. 13-50. Glass-reinforced plastic anchors of various types tested at VUIS. The surface of the anchor is smooth in the fixing section or threaded and provided with a base

Reinforced resin bars supplied, for example, by the Celtite group (France, England, USA) have a tensile strength almost identical to that of steel used for classical anchoring bolts (600 MPa),but a weight considerably lower (four times), and do not need anticorrosive protection. They can be placed reliably in the areas where cutting tools and machines are to be used because they can be cut easily and surely like a wooden rod. The bars can be prepared on site, adopting the procedures developed by the Spokane Mining Research Center in the USA [61]. The anchor consists of a glass-fibre rope, which is inserted into the borehole by means of special remote control apparatus (for reasons of safety), with resin and catalyst.

163

The latter components are kept in separate containers, and the rope is housed on a reel situated below the mixing and mounting head of the equipment. The fluid components are metred and pumped into the head, which is brought by the machine near to the roof of the excavation. The components are then mixed at a point close to the borehole mouth and the rope is pulled through the head while the plastic compound is pumped into the borehole. The boreholes may be of any length, and need not be straight. The resin in the hole sets within a few minutes and the machine can then proceed to the next borehole. This type of bolt is fastened to the rock along its entire length, is not prestressed, and is fully resistant to corrosion and the effects of shocks (Fig. 13-51). 13.2.4

Fixing of anchors by means of both cement and a mechanical base

Combined cement and base fixing is adopted only where short anchors (bolts) are used for the stabilization of rock. It provides the advantages of both mechanical and cementing methods, gives greater reliability of fixing in the rock, and affords the steel permanent protection against corrosion. The bolt can be prestressed immediately after installation, and its subsequent cementing to the rock along the entire length of the borehole prevents any losses of prestressing of bolt or rock, and enhances the resistance to extraction of the bolt. The method does not increase the load-bearing capacity substantially, because the resistance provided by the base and that attributable to the cohesion of the cement do not act simultaneously, but rather in succession, when the bolt is loaded. Only when the cohesion of the cement is overcome and fails does the mechanical base take the load. Either grout or polyester resin are used as the cement. The cement may be introduced into the borehole before or after installation of the bolt. Cement of less fluid consistency is placed in the borehole by hand in thin polyethylene wrappings which are ruptured when the bolt is inserted. Otherwise a hand pump (see Fig. 15-6), or a powered pump and grouting hose can be used (Fig. 13-52). A wedge-shaped base is most suitable, as this easily penetrates the cement to the bottom of the borehole. A convenient base for this purpose is formed by a cross-shaped cleft at the end of a shaped bar, and a cross-shaped wedge with conical termination (Fig. 13-53). The wedge bolts are inserted and rammed into the borehole already filled with cement, using a pneumatic pick of sufficient thrust (see Fig. 13-4). The borehole may be filled with cement after the insertion, fixing, and prestressing of the mechanical bolt, by injecting a more fluid mix under pressure (0.2 — 0.7 MPa) with the aid of an injecting tube and a pump. The borehole mouth must be sealed first of all. This procedure ensures a more thorough filling of the borehole and makes for the strengthening of any

164

fiberglass roving

polyester resin


Mechanical Expanding Anchorage

Fig. 13-51. Pumpable resin bolt of USBM

Vent Tube

Fig. 13-52. Dywidag combination rock bolt with threaded bar, expanding shell and injection tubes

165

Fig. 13-53. Kiruna type combination wedge bolt (Sweden) ä) — bolt bar with cross cleft, b) — cross wedging arrangement

jointed rock in its vicinity. If the rock is of compact structure without joints, allowance must be made for displacement of the air in the outer section of the borehole (see Fig. 13-52). A simple combination bolt can be made from thick-walled steel pipe, threaded at one end and with a cross-shaped cleft and conical wedge at the other (Fig. 13-54). After the pipe has been rammed on to the wedge resting at the borehole bottom, the filling of the borehole is carried out through this hollow tendon. To expedite the filling, the pipe is provided with lateral holes. A very effective type of combination bolt is offered by the American Williams Company (Fig. 13-55). Robust expanding shells of various lengths are used according to the type of rock, and these are fixed against the borehole walls by turning the bolting rod. The rod consists of a thick-walled pipe of high-tensile steel, its surface being specially shaped so as to interlock with the hardened cement and provide the possibility of simple extension of the rod by means of connecting pieces. Filling is carried out following prestressing of the bolt, either via the interior of the pipe or by means of a short plastic tube, always starting from the bottom of the borehole and proceeding outwards. A second tube allows air to escape and indicates the progress of the filling. For the sealing of the borehole mouth and the grouting itself, the manufacturer supplies special compounds based on quick-setting and expanding cements, but other materials can be used instead. 13.3 FIXING OF ANCHORS WITH ABUTTING BASES

Anchor roots which are to transmit large tensile forces especially in soils, should be designed as bases abutting on to load distribution structures built at, or sunk to, an appropriate depth of the ground (see Fig. 9-13). The load

& JA

^, !P,

VSSJ/M

JO ID 10

vt>>?i

PL

V//S>A

tinftttVAM

^fys/yjA

VSJSJX ±0v/Jssi

M 790

\

v/T;/i uoV;TS;A

I

Fig. 13-55. Williams combination rock bolt of hollow bar

HEX NUT

THRUST RlNO

EXPANSION

J^.'Sim

CROSS-SECTiON OF DE-AIR IN 1 RE-BAR'4

PfiiHG

v//s;jjj/j;jj//;;/.

////ί///Λ< // kO V;^"

^ j g ^ - - Φ - · - $ — $ ■ — Φ ~ —$ -

t-

Fig. 13-54. Combination tube bolt with conical spreading wedge (proposed by the Dept. of Geotechnology of the Czech Technical University, Prague)

section A-A'

167

distribution structures are usually sunk reinforced concrete parapet walls, trench walls, sunken wells (see Fig. 22-23), etc. In some cases it is possible to use the walls of construction pits made with rammed steel members or r.c. sheet piling (see Fig. 22-20). Where anchors are installed in newly made embankments, their bases can abut on to load-distributing slabs which are covered with fill as the work proceeds (Fig. 13-56).

Fig. 13-56. Anchoring of a precast quay wall using bearing plates embedded in the back-fill

Load distribution walls and slabs for the abutment of anchor bases must be placed a sufficient distance back from the anchored structure. This distance is determined by calculation (see Chapter 22), which must seek to ensure that there is sufficient resistance to loosening of the ground responsible for taking over the stresses from the load distribution structures and for securing the stability of the anchored structure. Reassurance must also be obtained, particularly for shallowly placed abutting structures, that these will not be deflected towards the surface of the ground when the anchors are prestressed, or when an increase in the load of the anchored structure occurs. Sometimes, for structural or architectural reasons, the anchor base is fixed directly in the anchored structure, while the head of the anchor abuts against a load-distributing wall (see Fig. 22-21). The use of abutting structures is uneconomical for anchor roots situated deep under the ground surface. Such roots are formed as bulbs by concreting the expanded end of the borehole in which the anchor end, spliced in the shape of a broom or provided with a strongly attached base, is inserted. In this fixing method the concrete-to-ground bond does not determine the fixing strength of the anchor, which is more a function of the diameter of the bulb and the compactness of the rock or soil. It is expedient to increase the soil compactness wherever technically possible, by injecting the concrete under high pressure, etc. The fixing of anchors by the expanded root (bulb) method is relatively little used in practice, but the results of research show that they

168

have a higher load-bearing capacity than anchors with long fixed roots. By using expanded roots, the length of both borehole and anchor can be reduced. Tests carried out by L. B. Underwood [214] on the St. Randall Dam site (USA) showed that anchor fixing by means of a root bulb is more reliable than methods of long anchor root based on the cohesion of cement. The tests were performed in a Cretaceous formation, with both horizontal and vertical boreholes. Seventy five bars were used, each 32 mm in diameter and equipped with a welded 114 mm diameter base. The diameter of the boreholes was 152 mm and that of the anchoring cavity 420 mm. The anchors were tested in 10 loading cycles, and only in the last cycle were they torn out. The greater reliability of anchors with bulbs was very evident, particularly under repeated or dynamic loadings. In Czechoslovakia a series of comparative tests was carried out on expanded and unexpanded roots fixed in sandy clay; the tests are described in Chapter 10 (see Fig. 10-28). Roots which are expanded induce by the tension of the anchor pressures and tensions in the ground, and the pattern of these stresses is similar to that resulting from a concentrated pressure at the ground surface. The limit strength of the rock or soil in the vicinity of an expanded anchor root (plate or bulb) is, however, several times greater than the limit strength at the surface. This fact makes for greater economy in the design and use of these anchors in practice. An anchor may be designed to take a considerable load without the necessity of making a large cavity for the bulb at the bottom of the borehole; it is not necessary to keep the stress loading on the ground around such a root within limits based on standards for surface loads. In practice it is possible, and indeed necessary in most cases, to assume that any ground failure takes place adjacent to the root, including the development of fissures near the base and at the bulb front, and plastic deformation of the entire ground in the vicinity of the anchor fixing. However, such changes in the rock or soil only appear at a certain depth below the ground surface, and extend over a very restricted area; further propagation is prevented by the strength of the ground around the damaged zone, there being a rapid decrease in stress with increasing distance from the fixing position. A failure in the ground next to the anchor base may be permitted without any risk of the uprooting of the anchor, provided the fixing depth is selected deep enough, according to the principles outlined in Chapter 10. This anchoring theory is supported by the results of many tests. In the tests on the Allt-na Lairige Dam site, described in Section 10.1, the stress created by the concentrated pressure at the front of the expanded base reached 188 MPa; similarly a stress of 132 MPa was measured in anchoring tests in the dolomitic limestones of the Cierny Väh river (see Section 28.4). Anchoring tests in dry, noncohesive soils in Nosice and

169

Sucany produced stresses of up to 53.5 MPa at the front of the anchor bulb, while in tests in loess (see Section 10.4) a value of 7.1 MPa was registered. The concentrated stress created directly at the abutment of the anchor base on the soil was not, however, the cause of failure in the latter case. The failure occurred in the loess, because its shear strength was exceeded along the lateral surface of an inverted cone with its apex at the anchor base [86]. 13.3.1

Design of anchors with root bulbs

The relationship between the load-bearing capacity of an anchor and the diameter, d, of an abutting disc base has been studied by Y. Barraud [9]. He found that there was a direct relationship between load-bearing capacity and the ratio h\d(h being the depth of the base in the ground). By experimenting with a mixture of fine sand and gypsum (γ = 1.5 to 1.6g/cm 3 ), he concluded that the stress was transferred from the anchor base to a body of soil of the shape of a truncated cone, with its apex at the anchor base. A total failure of the soil above the base would be expected only at shallow depths, where the ratio h\d does not exceed 10. Otherwise, failure of the soil occurs only in the immediate vicinity of the front of the anchor, whereupon the soil flows around it (local failure). T. H. Hanna [71], testing models in dry sand to the limit of overall soil failure, arrived at a minimum value of h\d = 13. According to H. Nendza [145], an exponential increase in load-bearing capacity was observed up to h\d — 14 for anchors with expanded bases fixed in medium and strongly compacted sand; at greater depths and with the same base diameters the relationship was linear. The transition between these two types of failure has also been studied in models by A. G. Müller and R. Haefeli [139]. They found that the angle of the uprooted cone of a sandy soil decreased from 60 to 10° with increasing depth of the anchor base. The transition from the one type of failure to the other has also been clearly demonstrated by L. N. Dzhioyev [46] in site tests on the load-bearing capacity of anchors fixed into clayey soil by means of spherical bulbs (described in Section 10.4). In both types of failure above the anchor base, the soil yields in the direction of minimum resistance. In the first type of failure, (general damage) the soil overburden in the shape of a truncated cone is lifted as a result of its small weight, and in the second type the weight of the overburden is too great, so that local damage due to soil compression takes place. When the anchor is pulled out to some extent, this soil compression is accompanied by a squeezing of the soil into the space vacated underneath the anchor base. The soil grains around a flat anchor plate move (according to tests performed by L. Hobst in the late fifties) along circular shear surfaces.

170

The type of soil damage that occurs above the anchor base is governed not only by the ratio h\d, but also, and more particularly, by the physical properties of the soil: its compactedness, moisture content, angle of internal friction, and cohesion, as demonstrated by experiments [71, 133, 145, 187, 104]. For a given set of soil parameters, it is always possible to calculate theoretically the depth at which local soil damage will change into general damage on the extraction of an anchor of given base diameter. A similar calculation can be made with respect to the diameter, d, at constant depth. The depth of an anchor with an expanded base should be such that under the maximum load-bearing capacity of the anchor, the transition from total failure to local failure cannot take place. In the design of anchors with root bulbs, the physical and mechanical properties of the soil must be considered first, and it is advisable to carry out load-bearing tests in this respect (see Section 10.3). The admissible stress, ac„ of the soil under coaxial pressure is determined from these load tests, and the minimum cross-sectional root area that will prevent soil flow around the root under the required anchor loading, JP, is calculated. Area of abutting base:

diameter of circular- section abutting base (bulb)

V ™cr It is clear from field tests that in good load-bearing ground (hard and soft rocks, dry non-cohesive soils), a bulb diameter 20 to 50 cm greater than the borehole diameter is usually sufficient to ensure that the critical load limit is not exceeded. In saturated and otherwise softer non-cohesive soils, and also in cohesive soils, the anchor bulb must be wider in order to achieve a load-bearing fixing. A diameter of 50 to 100 cm is generally required if the critical stress resistance of the soil, as ascertained in tests, is not to be overcome (see Section 10.4). The load-bearing capacity of anchors and their margin of safety, m, are determined by in situ tensile tests carried out prior to the erection of the structure (see Chapter 17). To calculate the bearing capacity of anchor foundation bases and inclined piles, the empirical method may be used as presented by A. S. Kananyan, M. J. Nikitenko, J. A. Sobolevskij and V. N. Sukhodoev in 1977 [104]. This method offers the possibility of plotting coincidence diagrams in terms of non-dimensional co-ordinates after conducting several model test at different model scales. The principle of approximate model testing is to create a model

171

of the geometrical relationships of the foundations, thus maintaining a dynamic similarity, the soil being the same in both the model and the actual structure. 13.3.2

Fixing of the tendon to the bulb

In anchors with terminal bulbs, there must be a connection between the tendon and the root. In the a circular load-distributing plate is fastened on to the or bundle of bars. The diameter of this plate should be that of the borehole (Fig. 13-57).

reliable and strong case of bar anchors root end of the bar, 5 to 10 mm less than

Fig. 13-57. Load-distributing circular base at the end of an anchor consisting of a bundle of three bars {Macalloy system)

The anchoring cavity, or, where appropriate, the expanded end of the borehole must be filled with concrete; the concrete bulb thus created extends the pathway that the flow of soil grains must follow when local failure of the soil occurs in the advance zone of an overloaded root. Where the load distribution plate is embedded directly in the soil, that is, where the borehole around the plate is back-filled with soil of the same structure, the anchor can be extracted by a much smaller force, since the flow of soil articles can occur from the front face of the plate to the rear along a relatively short pathway. However, in coarse-grained compact gravels, even this method of fixing provides considerable strength (see Fig. 10-17). The ends of cable anchors are spread out fan-wise to a length corresponding to the expanded region of the borehole; this guarantees thorough embedding of the individual cable wires in the concrete (Fig. 13-58) and a strong connection between the tendon and the concrete bulb. The wires also provide bulb reinforcement. When cable anchors are inserted into boreholes, the unwound and spread-out wires must be temporarily bound together with

172

Fig. 13-58. End of a cable prised apart into a broom shape

mild binding wire to allow the end of the cable to pass easily down to the anchoring cavity. After insertion, the binding wire is severed and removed, so that the cable wires spring out at the bottom of the borehole. Breaking cones are normally used to split the anchor end (Fig. 13-59). In the case of vertical boreholes, the breaking cones are pressed among the wires at the end of the anchor cable as the cable comes to rest under its own weight on the cavity bottom. In oblique and horizontal boreholes, the cones are pressed into the cable by means of a draw-bar which passes through the axial grout pipe of the cable, or by removable draw-bars placed at the sides of the inserted cable. An expanding bag has also been experimented with for spreading the wires, this being placed among the ends of the wires and attached to the grout pipe passing down the cable axis. The material of the bag should be thin enough to rupture when the binding holding the wires together is released. Then free passage of grout into the cavity is ensured. The fixing of the wires in the concrete of the cavity filling is improved by the gripping effect of the force R (arising from reaction Q) when the bulb is under tension (Fig. 13-60). If an anchorage experiences a force P, the latter is resisted both by adhesion and by the reaction, Q, of the rock, where

173

Fig. 13-59. Spreading cone for prising apart cable anchor ends {Hobst system)

2 cos (a — b is the angle of friction of steel on concrete. The anchoring cavities of vertical anchors are sometimes filled with cement mortar before the anchor is inserted, but usually the grouting is only carried out after the anchor has been inserted. In the latter case the borehole section that is to be filled prior to anchor stressing is delimited by a collar or a seal. The grout pipe must lead into the anchoring cavity, and a smaller pipe allowing air to escape from the sealed section of the borehole is inserted along with the grout pipe.

174

Fig. 13-60. Locking forces increasing the strength of an anchorage in rock

Fig. 13-61. Anchors used in Muda Dam 1 — pipe for asphalt grouting of the borehole, 2 — washer, 3 — grout sealing, 4 — spacing cylinder, 5 — anchor heads in which wires are fixed individually, 6 — sheet-steel mould, 7 — helical reinforcement, 8 — fixing of separate wires

In some cases the fixing efficiency of the cable end into the concrete bulb is further improved by the use of a base (a plate) attached to the anchor end. Sometimes these bases are made from short pieces of seamless steel pipe into which the looped ends of the wires are concreted in advance. All the above-mentioned methods of connecting the anchor to the bulb have been tested and shown to be reliable. An interesting method was used at the Muda Dam site in Malaysia [215]. Here, each wire (patented w ; re, 7 mm diameter) of the root section of the anchor, passed through the tapered hole of a small head, and the heads were secured at different root levels by small wedges. The system of small anchor heads created in this way was placed in a slightly tapered mould and embedded in concrete. When the concrete hardened, the root of the cable was inserted into the borehole (121 mm diameter) and was grouted in with a cement slurry (Fig. 13-61). The conical expansion of the root towards the borehole bottom provided a more efficient anchorage similar to that obtained with a bulb-shaped root.

175

The most suitable arrangement for an anchor bulb, and the most appropriate method of forming the anchoring cavity, largely depend on the properties of the rock or soil into which the anchorage is made. 13.3.3

Fixing of anchors in concrete structures by means of bulbs

In some cases there may be a need to anchor structures into preconstructed concrete foundations, or other massive structures. If such an anchorage is planned, small shafts (or pits) can be formed in the concrete structure to take the anchorage, and the bottoms of these shafts can be shaped as anchoring cavities. The right type of cavity is created by embedding in the concrete well reinforced (or prestressed) precast concrete components with tapering hollows, or by using seamless steel pipes (Fig. 13-62). For example, in the construction of a cofferdam at the Orlik Dam in Czechoslovakia, the anchor cables (of 4 MN load-bearing capacity) were fixed in steel pipes set in concrete, as shown in Fig. 13-62a. Φ 370

Fig. 13-62. Anchoring cavities formed in a concrete structure by embedding special steel bases in the concrete (proposed by VUIS) a)— type of base used in the 2ermanice and Orlik Water Projects (Czechoslovakia), b) — type of base used in the Bariri hydro-electric power station (Brazil), 1 — 300 mm dia. pipe of 16 mm wall thickness, 2 — load distribution cone lowered with anchor, 3 — corners filled with concrete, 4 — spreading cone welded to bottom of steel cavity, 5—seating plate

In the case of the additional cables installed in order to anchor the guide vanes of the Bariri hydroelectric power station in Brasil, the anchoring cavities were made by setting steel structures, as illustrated in Fig. 13-62b, in the concrete. At the bottom of each cavity a breaking cone was welded in position so as to spread the cable end.

176

13.3.4

Fixing of bulb anchors in rocks

Rock cavities are made by using various accessories attached to the boring equipment (see Section 14), or by using explosive charges. Excavation with explosive charges is quicker, but cannot be applied in all cases. When blasting a cavity, successive charges must be used since one heavy charge may cause major rock damage. An experimental cavity was made in flysch sandstones in a cutting; charges of 1.6 kg and 3.20 kg of Danubite 20 were inserted in a borehole of 80 mm diameter, and after serial blasting was carried out a cavity approximately 60 cm high and 40 cm in diameter was formed. The charges were fixed around the circumference of a cylinder 40 cm long. The length of the charge is determined by the length of anchoring cavity required, whilst the size of the charge used depends on the rock type and the required cavity diameter. A specially designed anchor fixing in rock using concrete bulbs was applied in the stabilization of a rock slope in cavernous limestones [228] (see Fig. 21-25). The anchoring boreholes, including the fixing sections, passed through large karst cavities which were impossible to fill. The tendon section destined for fixing was placed inside a linen bag, a grouting pipe was provided, and then the fixing section was fitted with a seal and inserted into the borehole. When the grouting was carried out, the bag, which had twice the diameter of the borehole, formed expanded sections of the root in the cavities (Fig. 13-63). On prestressing, these bulbs abutted against the rock and safely transferred the tensile forces of the anchors, which were prestressed to 600 kN, to the ground.

Fig. 13-63. Fixing of cable anchor in cavernous limestone by means of grout-filled linen bags (ace. to Zajic) 1 — anchoring borehole, 2 — bundle of patented wires in fixing section, 3 — spreader ring, 4 — locking sleeve, 5 — outline of linen bag after grouting, 6 — grout filling of bag, 7 — protective PVC tube of anchor tendon, 8 — pipe for grouting root, 9 — sealing at end of bag and at end of protective tube, K — karst cavities in limestone

13.3.5

Fixing of bulb anchors in non-cohesive soils

In compacted dry gravels and sands, the anchoring cavity is made with a reamer designed for widening boreholes. In saturated loose soils, making anchoring cavities is difficult. The walls of the boreholes tend to cave-in

177

when the casings are removed, and the use of a thick drilling fluid is necessary. An experimental anchoring bulb was made by setting off an explosive charge placed among the ends of the anchor wires. The experiment was carried out in a borehole at an angle of 52° to the horizontal, and 0.8 kg of Perunite 20 explosive was used with double detonating fuses and ignition wiring. The anchor cable was inserted into the borehole together with a polyethylene grouting pipe. Prior to blasting, the casing was pulled out 150 cm and the lower part of the borehole was filled with 20 litres of grout. Immediately after the blast, the root was thoroughly grouted with a further 60 litres of cement slurry, and in this way a bulb 100 cm long and maximum diameter 50 cm, was created (Fig. 13-64). A strength check confirmed that the cable steel had not been affected by the blast. When anchors are fixed in coarse non-cohesive soils (grain size over 5 mm), an anchor bulb of irregular shape can be formed simply by filling the unexpanded borehole with cement slurry under pressure, so that the slurry penetrates into the surrounding gravel. The cross-section of a root expanded in this way depends on the permeability of the ground and the grouting pressure applied. In sands, a substantial increase in root diameter can be achieved in this way (Fig. 13-65). An investigation has also been made of the type of material most suitable for the filling of the borehole above the root O.e. in the tendon section) in

Fig. 13-64. Anchor bulb created by the blast of an explosive charge placed among the ends of the wires of a cable anchor already grouted in saturated gravel and sand

178

non-cohesive soils. Laboratory and field tests indicated that concrete was less suitable for this purpose, and that compacted soil of similar composition to that of the borehole surrounds was much more advisable, since this increased the anchor fixing strength.

Fig. 13-65. Irregular shape of a bulb created by the high pressure grouting of a long anchor base (Weber system) in relatively impermeable sand (photo: Stump Bohr Co.)

An interesting method of fixing anchors in soil involving the use of a slanting pile has been published by the Soviet authors Nikitenko and others [148]. To anchor the stabilizing cables of electric pylons in soil, r.c. piles 56 cm in diameter and 3 m long were rammed or inserted into oblique boreholes; the piles were provided with a tip up tie-bar equipped with a removable cutting edge (Fig. 13-66). After the pile has been installed,

w_

77J777777777777777777777777Z&77777.

Fig. 13-66. Anchor root formed by slanting pile [148] and hinged draw bar 1 — pile axis, 2 — hinged draw bar (diameter 36 mm, length 190 cm) with removable cutting edge, 3 — steel sleeve by which draw bar is attached to pile, 4 — draw bar in final position after having been fixed to a cable from the pylon, 5 — sand and gravel filling of the borehole after installation of the pile

179

the tie-bar is drawn laterally so that it cuts through the soil, until it finally reaches the required position whereupon it is connected to a cable from the pylon. In tensile tests this anchorage registered a load-bearing capacity of 150 to 160 kN in coarse sand, 100-110 kN in fine sand, and 70—&QkN in loamy sand. The advantage of this method is that it saves time and excavation work. 13.3.6

Fixing of bulb anchors in cohesive soils

Bulb anchors are the best for cohesive soils, because in such soils the advantage of bulb fixings over unexpanded root fixings is greatest, and also because cavities made in cohesive soils do not tend to collapse. The cavities at the ends of the boreholes are made with special drilling tools fitted with reaming knives, or by exploding small charges in the boreholes (Figs. 13-67 and 68). Drilling poses the problem of how to remove the loosened material from the cavity as drilling proceeds (see Section 14.2.5). Borehole expansion

Fig. 13-67. Anchor bulb formed by concreting a cavity bored in loam

Fig. 13-68. Anchor bulb created by concreting a cavity formed by blasting. The end of the wire cable was prised apart by a cone

180

by blasting is only suitable for deeper anchors, since blasting close to the surface (i.e. less than about 5 m) causes widespread damage in the surrounding soil with the result that the fixing strength of the anchor is impaired. A view of such a cavity formed by blasting is shown in Fig. 14-15. The load-bearing capacity of bulb anchors depends largely on the area of cross-section of the anchor bulb, and cavities should therefore have the maximum possible diameter. For fixing the tendons of large cable anchors in cavities formed either by blasting or by mechanical means, the spliced end of the cable is spread, as the cable approaches full insertion, by a cone resting at the bottom of the cavity; this procedure is followed even in cohesive soils. The cavity is then grouted. The transverse stresses occurring at the point where the tendon merges with the root are countered by a strong cable bandage, formed either by leaving a part of the borehole casing in the lower section of the borehole, or by inserting a helical reinforcing coil into the upper part of the cavity. Such measures are not necessary for bar anchors with simple anchoring bases. A different type of expanded root anchor fixing has been introduced by some British companies. Instead of creating one large cavity at the end of

Fig. 13-69. Excavated root showing two bells of a Fondedile anchor

181

the borehole, special drilling and reaming equipment is used to make several successive borehole expansions of two or four times the shaft diameter, each with the shape of a truncated cone or bell (Fig. 13-69). The Universal Anchorage Co. achieved the following service load-bearing capacities of anchors by this method: 0.25 MN in a clayey soil of cohesion c = 0.1 MPa; 0.50 MN in a gravel and sand soil; 1 to 4 MN in rock, according to rock type. The Fondedile Foundations Ltd. prepared for its Multibell anchor system a guide line of an approximate ultimate load-bearing capacity as shown in Table 13-V. A safety factor of 3 is generally applied for permanent, and of 1.5 to 2 for temporary, anchorage. TABLE 13-V. Ultimate load capacity of Fondedile Multibell anchors Number of bells

2 3 4 5 6 7

Bell Bell Bell Bell Bell Bell

anchors anchors anchors anchors anchors anchors

Shear strength of clay 120 kN/m2 160 kN/m2

200 kN/m2

400 kN 590 kN 790 kN 980 kN 1,180 kN 1,380 kN

660 kN 980 kN 1,310 kN 1,640 kN 1,970 kN 2,300 kN

520 kN 790 kN 1,050 kN 1,310 kN 1,540 kN 1,830 kN

Test results on these Multibell anchors in typical London clay are presented in Fig. 13-70. A root in service having seven bells as shown in Fig. 23-17. The excavated root of a Universal anchorage Co. bar anchor is shown in Fig. 13-71. The increased load-bearing capacity of anchors with roots consisting of a succession of bulbs results from the greater area of contact developed between root and soil when the anchor is pulled out. By increasing this area, it is possible to increase the loading force on the anchor before the ultimate stress, acr9 is reached, this stress being critical for the security of the anchor against extraction of the root by cutting through the soil (see Section 10.4). This stress is the governing factor in the determination of the fixing strength in soils of lower strength. An increase in the number of bulbs in an expanded root has, at least in the initial stages of prestressing, about the same effect as an increase in bulb cross-section. Forming a greater number of smaller cavities is technically easier than creating a single large cavity at the borehole end.

182

Fig. 13-70. Results of load tests on Fondedile Multibell anchors in typical London clay

7 bell anchor. 'c'valueof jJay-MktUffl2 5 beJUa/fcbor 2 c « MkH/m 3be//anchor c- I70kfl//n2

10

20 30 ¥0 50 Extension in millimetres

Fig. 13-71. Two bells of an Universal Anchorage Co-bar anchor

The studies of L. Hobst have shown that the resistance of the soil medium as a whole against extraction of the anchor is not increased substantially when multiple bulbs are used in place of single larger bulbs. The conical shear surface develops only above the upper bulb of the series. The soil which is displaced by the movement of the lower bulbs, is transferred into the space vacated by the upper bulbs and does not contribute substantially to the fixing strength of the anchor. The manufacturers of Multibell anchors usually recommend to design these anchors under the assumption that failure of clay occurs along the cylinder defined by the tangents of the bells and containing all the bells. Thus if the shear strength of the clay is known, the number of bells and their depth may be determined by simple calculation.

Chapter 14 D R I L L I N G OF A N C H O R

BOREHOLES

The drilling of the boreholes is usually the costliest operation in anchoring, and it is often, therefore, the determining economic factor in decisions about whether to employ anchoring at all. Clearly the most efficient drilling methods must be selected, and the time schedule of the drilling operations must be carefully estimated. Two basic types of drilling can be distinguished, namely, drilling for short anchors (bolts) of small load, and drilling for long anchors transmitting large tensile forces.

14.1 SHORT BOREHOLES OF SMALL DIAMETER

For short bDlts, that is, for boreholes up to 3 to 4 m in length and 45 mm in diameter, ordinary hand-operated percussion drills are adequate. To obtain the necessary thrust and correct guidance, particularly in back holes directed upward, pneumatic props or mechanical braces between the roof and the floor of the gallery must be used, as well as a guide along which the hammer drill moves (Fig. 14-1). In other cases, multi-purpose pneumatic hammer drills are normally used, these machines being economic, timewise, in all anchoring applications. The Swedish Atlas Copco pneumatic hammer, type Falcon BD 46, weighing 35 kg, drills at a considerable rate (20 to 30 cm/min). It also drives the bolt bar into the borehole and tightens the nuts with a moment of 343 Nm; -usin^special adapters. The English Victor drilling hammer is similar, with the drilling machine mounted on a guide column braced between the roof and gallery floor (Fig. 14-2). The hammer is available with a pneumatic drive or an electric drive, and is able to remove the dulst from the borehole during dry drilling. When larger underground caverns are strengthened with anchors, the boreholes, even those for short bolts, are drilled with the efficient and mobile Jumbo machines, which have one or two drilling arms and are operated from a control panel on the carriage (Fig. 14-3). Some of these machines (e.g. Ingersoil—Rand, Secoma, Atlas Copco, Tamrock, Böhler, Alimak Montabert) are specially designed for anchoring operations, and are adapted and automated for driling boreholes, placing the bar anchors into the boreholes (together with resin, if used), and prestressing the anchors to the

184

U'sA

f l f t l * ! ^..V.

Fig. 14-1. Drilling hammer on light thrusting equirment used in Czechoslovakia (photo Osan)

force selected by the operator (see also Fig. 20-21 and Chapter 20). For a smaller volume of work a light mobile drill (Fig. 14-4) may be very useful.

185

Fig. 14-2. Multi-purpose drilling hammer of the English Victor Wallsend Co.

Fig. 14-3. Jumbo drilling equipment of Ingersoll-Rand being used for anchoring in an underground opening

Fig. 14-4. Alimak BT 121a light, self-propelled, compressed-air-driven carrier for mechanized drilling

186 14.2 BOREHOLES OF LARGE DIAMETER AND LENGTH

Anchors of high load-bearing capacity generally require long boreholes (5 to 50 m) of large diameter (60 to 150 mm). However, the maximum volume that is to be handled by drilling restricts the size of boreholes to less than 100 mm diameter and 25 m in length. 14.2.1 Suitable drilling methods For long boreholes, percussion drilling, rotary drilling, or the two combined may be used. A rotary drill (Fig. 14-5) transmits two basic actions to the rock through the drill rod and circular three-cone (roller) or auger bit

Fig. 14-5. Rotary drilling set (Hausherr Co.) in use for making slanting boreholes for anchoring the walls of construction pits (documentation of Soletanche Co.)

(Fig. 14-6), viz. axial thrust and rotational torque. Percussion drills (Fig. 14-7) penetrate the rock by the action of repeated impulse blows, usually from a chisel or wedgeshaped bit with hammer and drill rods. The torque, rotational speed, and thrust requirements are significantly lower for rotary —percussion systems than they are for rotary systems. The flushing media most commonly used to remove particles of rock from the drill bit are air, water or "mud", the latter usually being a suspension of bentonite in water.

187

Fig. 14-6. Wirth B-O rotary drilling set with auger bit, in use in loamy deposits (photo Geotest)

, friftL

JP^ 1flPw*U

■!1 B I »BidNiliii^iiiilSsii!«!

»jBHKHBBiii!

e^t"

Fig. 14-7. Tamrock percussion drilling jumbo with three booms for use in underground excavations

188

Suitable drilling equipment must be selected with regard to the type and quality of the rock, the diameter and length of the borehole, the accessibility of the anchoring site, the type of flushing medium to be used, the anchor type, and the required drilling rate of the machine. The rock type and the dimensions of the boreholes are the most important factors in most cases. A guide to the most suitable drilling method for a particular type of rock and diameter of the borehole is given by the diagram in Fig. 14-8 compiled

3^ Co

S-

<-o

ry

rota

^ }

0

\D-T-H

300

1 J

^rt -\200\ 11 ^ 1 A

Co

1

/e

percuss] I,

^ ^

1

1

... _..

■ \

\

Fig. 14-8. Preferred methods of drilling according to class of rock and hole diameter [129] (D — T— H = down-the-hole hammer)

by McGregor [129]. In cohesive plastic soils, rotary drilling with an auger bit without flushing is most appropriate; in loose soils and friable rocks, rotary drilling with a roller bit and water, or better still, a thick flush, is more efficient. In strong rocks, percussion drilling with air flushing is the most convenient method where the borehole diameter is small, whereas percussion drilling with a down-the-hole hammer is preferable for larger diameters. Rotary drilling with circular hard-faced bits are used in softer rocks, and diamond bits are required for hard rocks and very large diameters, whether in hard or soft rock. Water is always used as the flushing medium. The advantage of the last-mentioned drilling method is that it yields cores which provide information about the quality of the rock along the entire length of the borehole. The disadvantage is that the walls of the borehole are very smooth with the result that the grout-rock bond is lower (Fig. 14-9). For boreholes which traverse various types of rock, McGregor recommends that only the diameter and length be considered. He provides a diagram (Fig. 14-10) as an aid to selecting the most appropriate drilling method. Some authors (e.g. Parker [163]) consider percussion drilling by light wagon drills to be the most effective in all types of rock and soil with the exception of soft ground, provided that the borehole diameter is less than 100 mm and its length does not exceed 50 m. When the borehole is to be drilled through a deep soil into strong underlying rock, the Duplex combined drilling system developed by Atlas Copco of Sweden can be used to advantage. This drilling machine can work with

189

mm -€

.WE I

.*#*^v*1

's?

Fig. 14-9. Drilling of boreholes using core drilling equipment supplied by Stump Bohr AG of Switzerland; the placement of the anchors in casings is shown

|00 200 300

percussive-rotary -| drills

__L

rotary JL

4-

diamond drills

L

drills J

L

0.3 1.5 3.0 7.5 15 30 60 150 300 depth (m)

Fig. 14-10. Preferred drilling methods in mixed strata [129]

a percussive drilling rod or with a rotary external annular bit together or with both of these if required of the borehole in loose soil or weathered soft reck is reck the drilling proceeds with the central percussive 14.2.2

casing fitted with an (Fig. 14-11). The part cased, while in strong bit alone.

Work rate of the drilling equipment and drillability of rock

The work rate of a drilling machine is measured in terms of the length of a borehole of predetermined diameter drilled per unit time. Usually this is determined in a direct test on rock of a defined type. Percussion machines

190 Fig. 14-11. Double drilling system using a drill rod set and casing, introduced by the Sandvik-Coromant Co. of Atlas Copco, Sweden

of the wagon drill type, drilling 100 mm boreholes in strong rocks, with air flushing, attain rates of 5 to 15 m per hour. Rotary core drilling machines work at half this rate at best. The drilling rate depends on a number of factors including the condition of the machine and drill bits, the flushing method employed, the air pressure (in pneumatically driven machines), the torque developed (in rotary machines), and the rock type and borehole diameter. The power output needed to drill a borehole in a particular rock is referred to as the drillability of that rock. This characteristic is largely dependent upon the hardness of the rock; usually, the harder the rock the more difficult is the drilling. Drillability, however, also depends on the mineral composition, the strength of the rock, the grain size, porosity, stratification, and density and direction of the joints, etc. No generally accepted procedure for determining the drillability of rocks has yet been worked out. The best known test for establishing the coeffic;ent of rock strength was published by Protodyakonov [169] and subsequently modified by the U.S. Bureau of Mines (1968). Basically it involves fracturing rock samples by an impact of known value, and then weighing the broken pieces. 14.2.3

Flushing

The flushing method used may markedly influence both the rate of drilling and the quality of the borehole. Air flushing is the most commonly used method for percussion drilling and roller bits. It is highly efficient in dry rocks, and may also be used in moist situations if a sufficient quantity

191

of air is available, although its efficiency does not differ greatly from water flushing in the latter case. Water flushing is used mostly in rt>tary core drilling; it results in clean boreholes and makes for a good grout-rock bond, even where percussion drilling is employed in moist rock or saturated noncohesive soil. On the other hand, if water flushing is used in clays, marls, and other cohesive soils, it must be reduced to a minimum because the flushing causes a deterioration of the mechanical properties of these soils and reduces the shear resistance at the anchor root surface. The design length of the anchor borehole must be extended by 30 to 70 cm to create a "sump" for the debris and mud which cannot be removed by the flushing. When drilling is terminated, it is important that flushing of the borehole is continued from the bottom up for at least 10 minutes [120]. 14.2.4

The anchoring site

The accessibility and situation of the anchoring site may be the decisive factors in the selection of the drilling method and type of drilling machine. Considering that the anchors are usually arranged in parallel lines, a drilling machine mounted on a wheeled undercarriage is an advantage (see Figs. 14-4, 12), the more so if it can be power driven from one anchor position to another. Drilling machines mounted on crawler tracks need to be used on the uneven and unreliable floors of foundation pits (see Figs. 14-5, 6). Machines for drilling anchor boreholes must be able to drill at any angle from the horizontal to the vertical up to a height of 2 m above the base. Percussion drilling machines present difficulties in built-up areas because of the noise produced. The admissible noise limit in urban areas is 70 dBA at a distance of 15 m from the source of the noise. The use of percussion drilling machines with air flushing is forbidden in subterranean rooms on account of the great quantity of dust which is produced. In both types of situation, percussion machines are being replaced by rotary drilling sets with water flushing. Frequently the need to drill boreholes from a scaffolding arises. Provided the location is accessible, and a heavy, rigid scaffolding can be constructed with a load-bearing capacity of at least 5 kN/m 2 , then equipment of the wagon drill type weighing about 1,000 kg can be mounted and used from this scaffolding (Fig. 14-12). The horizontal forces created by these machines are not very great and therefore can be withstood easily by the scaffolding and mounted working platform. Drilling becomes more complicated in places with poor accessibility such as steep rocky slopes, where only light or possibly suspended scaffolding can be erected. In these situations, only the lightest drilling sets can be used, and they have to be transported and assembled by hand on the scaffolding. In such cases, light, simple, electrically

192

Fig. 14-12. Böhler ET 11/35 percussion drilling machine, placed on a scaffolding and making anchoring boreholes of 70 mm in diameter and 25 m long in limestone

or pneumatically driven rotary sets are convenient, as they do not transfer any major vibration to the scaffold structure; their drilling rate, however, is relatively low. Fig. 14-13 shows the operation of such a set on a light scaffolding suspended over a railway line. An adapted mine core drilling machine with pneumatic drive was used; this older machine weighed 535 kg and drilled slightly inclined boreholes 61 mm in diameter and 20 m long in limestone at an average rate of 0.5 m/hour. Nowadays, light drilling machines with a much higher work rate are available on the market. For drilling short anchor boreholes from a scaffolding, the more powerful of the pneumatic hand hammers can also be used (Fig. 14-14). 14.2.5

Expanding of anchor boreholes

Boreholes are sometimes expanded at the innermost end in order to increase the load-bearing capacity of the anchor. This expansion is carried out by means of a special mechanical boring tool, or by blasting a small quantity of explosive in the borehole.

193 Fig. 14-13. Drilling boreholes on a steep rock slope near the Tetin ruin (Czechoslovakia) a) — scaffolding on the steep slope above the railway line,

b) — drilling of boreholes from an assembled platform. The rock surface is covered with synthetic netting

194

Fig. 14-14. Drilling short anchoring boreholes from a scaffolding on a rock slope using aPermon VK 22 light drilling machine (diameter of boreholes 45 mm, length 7 m; phonolite)

A number of tools of varying complexity and efficiency have been developed for expanding a borehole to four times the original shaft diameter, Thus, for example, the French Soletanche Company uses equipment shaped like a core barrel, which, depending on the sense of rotation, either extends or retracts the longitudinal reaming knives. The device cuts a barrel-shaped cavity. The Calweld company supplies a very similar tool for producing larger borehole diameters. There is one problem common to all reaming equipment, i.e. that of the continuous extraction of the reamed material from the borehole. Unless the work is done very carefully, the borehole may sometimes become overfilled, so that the jaws or knives cannot be retracted and the tool becomes jammed in the borehole. The extending knife (wing) of the West Ger-

195

man Klemm Company disintegrates the material of the borehole wall to such extent that it can be removed by water flushing. However, even with this equipment the operation is not always entirely successful, the actual diameter of the expanded region sometimes being smaller than the diameter potentially attainable with maximum projection of the knife. This reduced performance occurs particularly in the case of horizontal boreholes. The British company Fondedile uses for its Multibell system a special tool which reams the borehole, drilled in clay, to the shape of several bells at the same time (see Fig. 13-69). The reaming tool consists of a number of hinged blades which open in succession to form a series of quadrilaterals equal to the number of bells required, while the spoil cut by the blades is brought to the surface by direct circulation of flushing water. Thus all the bells are formed in one operation. Mechanical methods of reaming boreholes are suitable for compact and cohesive soils where the walls of the newly made cavities are to remain intact for some time. Borehole expansion can be achieved more efficient by using small explosive charges, this method being applicable to all ground types, even in noncohesive soils below the ground water level (see Fig. 13-64) where mechanically reamed boreholes would cave in. The diameter of the extended part of the borehole made in this way is usually larger and more irregular compared with that made by a mechanical reaming device (Fig. 14-15). The blasting can be effected either in an open borehole prior to insertion of the anchor, or after the borehole has been filled with grout and the anchor inserted. In the latter case the charges are placed among the reinforcing elements. This method has proved to be effective in tests. However, borehole expansion by blasting requires considerable experience in deciding on the size of charge that is most suitable. Misuse of explosives by unqualified persons can damage the ground and the anchorage, greatly reducing the load-bearing capacity of the anchor. 14.2.6

Defects of anchor boreholes

The most frequent defects of anchor boreholes arise from incorrect positioning and alignment of the borehole, and deviation or curvature of the axis away from the intended direction and inclination. Correct alignment of a borehole is a matter of careful preparation and the use of appropriate surveying instruments. Straightness of the borehole direction depends on the drilling set and the drilling technology. Boreholes become curved when the drilling rods are too slender, when excessive thrust is applied, and when there is a tendency for the bit to follow joints or other planar rock features which cross the borehole obliquely. Rotary core diilling

196

Fig. 14-15. View of a cavity made in a 137 mm dia. borehole in sandy clay by a charge of 150 g Semtex 186. Note the cracks in the cavity face

is particularly prone to curving in long boreholes. Deviation from the straight may be as much as 1 : 10 in a faulty drilling operation. Boreholes drilled by the percussion down-the-hole method show the least tendency to curving. In horizontal and inclined boreholes the weight of the drill rods presses the rods against the lower side of the borehole, nearly always causing a slight upward curving of the borehole. If the curvature of the borehole is sufficiently great, the drill rods damage the borehole walls and cause fragments of rock to fall into the borehole. Re-straightening of a borehole during the course of the drilling is very difficult. In checking the direction of an anchor borehole, a deviation of up to 1° (2 per cent) [120] may be admitted. 14.2.7

Permeability of anchor boreholes

Permeability testing of anchor boreholes is often demanded after completion of the drilling operations. The most suitable type of grout for the fixing of the anchor and its anticorrosive protection, the volume of grout required, and the most appropriate grouting pressure may be reliably established from the loss of water in the permeability test. If the losses are considerable, the permeable ground must be sealed first of all with a thick grout, and then the

197

borehole must be rebored after 24 hours and a further permeability test carried out (see Chapter 18). Pressure tests in strong rock are carried out in borehole sections from 1 to 5 m long, using one or two seals and beginning at the bottom of the borehole. The pressures used in the test must not exceed values which might damage the rock of the borehole walls, or open the joints. The upper pressure limit can be taken as twice that of the overburden lying over the section of the borehole in which the test is carried out.

Chapter 15 GROUTS AND METHODS OF G R O U T I N G A N C H O R B O R E H O L E S

Most anchors are either bonded with the ground, or are protected by a fluid cement mix (slurry), which on hardening forms a strong filling of the borehole, and may also strengthen the immediate rock or soil. The strength of the bond and the effectiveness of the anticorrosive protection depend very much on the composition of the mix, the thoroughness of its preparation, and the method used to forward it into the anchor borehole.

15.1 COMPOSITION OF CEMENT G R O U T S

These grouts are usually prepared from ordinary good quality Portland cement and clean water with or without the addition of fine-grained sand. The Portland cement should be fresh (not older than one month) and high grade (at least 300). The quality of cement deteriorates with age and is reduced by damp or over-hot storage, or storage in large quantities. The mixing water should be of the right quality. Water with a high content of sulphate (more than 0.1 per cent), chloride (more than 0.5 per cent), sugars, or suspended organic matter, is not suitable. Generally, any water which is suitable for drinking can be used for cement [120]. The sand used in the grout mixture should be fine (up to 2 mm grain size) and clean, without any loamy content. Sand is added to the cement in a weight ratio of 1 : 1 to 1 : 2. The weight ratio of water and cement is particularly important for the quality of a cement grout. Excess water results in bleeding of the mix and low strength, as well as greater shrinkage and lower durability of the hardened grout. The strength of hardened grout in relation to the water/ cement ratio is shown in Fig. 15-1 and 15-2. Experience shows that the most suitable water/cement ratio for grouts used in the fixing of anchors lies within the range 0.4 to 0.45. At these ratios, the grout is still sufficiently fluid for pumping, and penetrates easily into small openings and pores; the hardened grout is sufficiently strong and waterproof, showing little shrinkage. Additives can be put in grouts to accelerate or retard setting, to prevent shrinkage of the grout in the course of setting, to induce expansion, to increase the fluidity of the mix at low water/cement ratios, to prevent bleeding

199

*>20 £ 10

I»

'ß3 OU 0.5 0.6 0.7 0.8 0.9 10 water/cement ratio

Fig. 15-1. Effect of water content on the compressive strength of grout [120] Fig. 15-2. Gain in strength with time of ordinary Portland grouts at various water/ cement ratios (7 — 0.40, 2 — 0.45, 3 — 0.50, 4 — 0.60) [120]

2

3 4 5 7 10 time [days]

n

2d

of the mix, etc. The additives and respective dosages required to bring about the above-mentioned effects, are shown in Table 15-1. When additives are used, good quality grout is of particular importance. A combination of several types of additive with the intention of obtaining a combined effect is not recommended. It may be stated generally that experience in the use of grout additives is still very limited. Additives should therefore be used only when absolutely necessary, and then very careful preparation and control of the grout becomes essential. Additives for grouts are currently available under special trade marks; e.g. VSL Companies use Sica-Intracrete additives for anchoring grouts [122]. Apart from ordinary Portland cements, special sulphate-resistant and TABLE 15-1 Additives for grouts [120] Additive

Active chemical or mineral constituent

Optimum dosage (% of cement by weight)

Remarks

Accelerator

Calcium Chloride

1-2%

Retarder

0.2—0.5 %

Expander Anti-bleed

Calcium Lignosulphonate Aluminium powder Cellulose Ether

0.005—0.02 % 0.2—0.3 %

Fluidifier

Bentonite

2-3%

Accelerates setting and hardening Retards setting and increases fluidity up to 15 % expansion Equivalent to 0.5 % of mixing water Also acts as anti-bleed

200

rapid hardening cements are sometimes used. High Alumina Cement is a well known example; in the fresh state it attains 50 per cent of its final strength within 24 hours of mixing. This cement is recommended only for short-term test anchors, in view of the larger quantity of water needed to ensure fluidity of the grout, and the long-term volume instability of concretes made of this cement. The strength of the hardened grout is also very important. Strength depends not only on the type of cement used, but also on the hardening time. Usually a compressive strength of about 30 N/mm 2 (30 MPa) developing within 7 days of hardening is the minimum required. The effect of the water/ cement ratio and hardening time on the compressive strength of grout prepared from ordinary Portland cement is shown in Fig. 15-2.

15.2 PREPARATION OF CEMENT GROUTS

The following basic rules according to Littlejohn and Bruce [120] must be observed in preparing good quality cement grout: a"i the cement and the filler (sand), if applicable, must be measured by weight; b) water in the quantity required for the most suitable water/cement ratio must be transferred to the mixer before the cement (and fillers); c) any additive should be placed in the mixer, carefully measured out, during the latter half of the mixing time; d) although the mixing time depends on the type of mixer, it should not be less than 2 minutes. Mixing grout by hand should be avoided. The mixers used for the preparation of grout must ensure perfect intermixing of the cement, and must be able to produce grout of uniform consistency. This can best be achieved in small, rapidly rotating mixers with speeds of 1,500 —2,000 rpm. The Swedish Cemag mixer of the Atlas Copco Company (capacity 175 litres, weight 255 kg, Fig. 15-3), or the English Colcrete mixer, satisfy these requirements. After mixing the grout should be stored in a special tank and slowly agitated before use. Such a storage tank is shown in Fig. 15-3. The mixers, pumps, and delivery pipes should be kept scrupulously clean to ensure optimum output and smooth operation. The equipment must be attended to throughout the operation in case obstructions occur in the filters, the delivery outlets of the mixing vessels, or the bends and couplings of the pipes. Delivery pipes with an inside diameter of 12 to 35 mm are used; these are made of flexible material rather than steel, since this allows the position of any obstruction to be found without delay.

201

Fig. 15-3. Complete grouting equipment of the Swedish Atlas Copco Company in service in a gallery (from left: ZHS pump, Cemag 350 mixer, Cemix 75 storage tank and batching vessel for water)

15.3 F O R W A R D I N G OF C E M E N T MIXES INTO BOREHOLES

The method of forwarding the cement mix into the borehole must guarantee complete envelopment of the anchor and filling of the borehole; if necessary, it must also result in a strengthening of the surrounding ground. 15.3.1

Hand-filling of boreholes

Preparation and forwarding of the cement mix by hand is the simplest method, but this is only suitable for small quantities of grout placed in short boreholes free from the presence of ground water. Otherwise, the preferred method depends on the quantity of grout that is to be handled, the consistency of the mix, the borehole direction, and the type of bar anchor (bolt) that is to be grouted. The best known method is the Perfomethod; two longitudinal halves of a sheet metal tube, perforated with 1 cm holes, are filled by hand with a very thick grout. The two parts of the tube are then bound with a wire and inserted into the borehole (Fig. 15-4). The action of driving in the bolt bar causes the grout to be extruded into the space between the tube and the rock; after

202

Fig. 15-4. Placing of grout in a borehole by means of a perforated sheet metal tube (Perfomethod) 1 — sheet metal tube in halves, 2 — filling of several dismantled tubes with grout, 3 — driving of the anchor bar into the tube, thus forcing the grout into the borehole, 4 — cross-section of the filled borehole

borthok

Fig. 15-5. The grout contained in a tube of thin plastic material is pushed by the bolt into the borehole

203

hardening of the grout, all these components aie bonded together. In other cases, a tube made of fine-meshed wire netting has been used [105], the grout being pressed into it quickly and easily from the sides. Elastic tubes of polyamide, which are very cheap, have been used successfully for the transport of fluid grout. Bsing smooth and pliable, these tubes circumvent borehole irregularities more easily than rigid tubes made from metal sheet or glass. They are supplied in a range of diameters and can be used for fillings up to a length of 80 cm, although several lengths can be inserted one after another. When the filled tube has been pushed to the end of the borehole by the bolt (Fig. 15-5), the wrapping bursts under no more than hand pressure, the grout escapes, and the wrapping is pushed at the tip of the bar to the borehole bottom. A conical collar of soft rubber, previously slipped on to the bolt bar, prevents backflow of the thin grout as the bar is inserted, delimits the fixing section, centres the bar in the borehole, and holds the bolt and grout, even in an upward-directed roof borehole, until the grout becomes hard [228]. Another simple method is based on the bicycle pump principle. By moving a rubber piston in a plastic tube, fluid grout is drawn in from the supply vessel and then pumped into the borehole (Fig. 15-6). Then the bar fitted with a sealing collar is inserted. However, this kind of pump is only practical for filling short boreholes up to 2 metres long.

Fig. 15-6. Forwarding grout into a borehole by means of a plastic hand-operated air-pump

204

A very efficient method of forwarding thin grout into boreholes uses a hose and grout pipe connected to a grouting tank which is subjected to a low air pressure. This method is particularly suitable where a larger quantity of grout is needed for filling boreholes completely. The method used to grout boreholes is also determined by local site conditions, such as the availability of equipment at the site, the purpose of the anchorage, the anchor length, and the number and position of the bolts. Usually from 300 to 1,000 cm 3 of grout must be introduced into each borehole for anchoring the ends of short prestressed bolts. When calculating the quantity of grout required for a particular borehole length, that is, the space in the borehole around the bar in a given section that is to be filled, approximately 33 per cent, must be added, according to experience gained both in Czechoslovakia and elsewhere. This extra quantity is needed on account of irregularities of the borehole walls compared with a true cylinder, and allows for grout losses during handling (see Table 13-111, Section 13.2.2.1). 15.3.2

Grouting under pressure

Cement grout is forwarded into long boreholes under pressure, using mechanical piston pumps with electric or pneumatic drive. These pumps operate at pressures from 0.1 to 12.0 MPa; light weight and small dimensions are preferred to high pumping output. The grouting of anchors usually does not involve large quantities of mix; however, high quality of the grout and short forwarding distances in pipes of small diameter are prerequisites to successful grouting. Consequently, light, mobile grouting equipment is the most convenient. The Swedish grouting pumps supplied by Atlas Copco deliver 50 litres of mix per minute (piston diameter, 110 mm) under pressures of up to 7 MPa, or 30 litres per minute (piston diameter, 80 mm), under pressures of up to 12 MPa. The weight of these pumps including the driving unit is about-350'kg. The pumps of other renowned manufacturers, such as Wirth, Häny, Clivio, Meynadier, and others, have similar specifications. Screwtype pumps such as the Colmono and Moyno, are used in England and the USA. All the specialized equipment for the grouting of anchors is kept at one location at the site; the equipment includes, besides a pump and a high-speed mixer, a special low-speed mixing tank for grout storage. The latter is a vessel in which the mix is gently kept in motion until it is required for use. The storage tank also serves as a batch unit in the control of grout production. At sites requiring large volumes of grout, a high capacity facility for grout preparation is established at one location, employing a number of mixers and pumps. The grout is then distributed to various points through large diameter pipes.

205

For a smaller volume of work, and for places of difficult access including underground galleries, grout-making facilities assembled from light-weight units (see Fig. 15-3) so as to be easily movable, are more convenient. A complete Many grouting set-up mounted on a steel frame is shown in Fig. 15-7, and another type mounted on wheels is shown in Fig. 15-8. The special light-weight portable Spedel mixers and pumps weigh only 15 kg. They are driven pneumatically, producing a pressure of up to 2 MPa at which the output is 20 litres per minute. The grout is forwarded from the pump to the borehole through £ pressure rubber hose, and for the greater part of the length of the borehole, it is passed through a plastic pipe of inside diameter 12 to 25 mm, depending on the

206

Fig. 15-8. Mobile grouting set of Geoindustria Praha

operating pressure and the grouting method used. The grout pipe leading to the root is located within the anchor tendon or alongside the insulating wrapping, as far as the remote end of the anchor root. This pipe is usually short of the borehole bottom by 150 mm, and its end is protected from clogging while it is being inserted into the borehole together with the anchor. The pipe is attached to the anchor with adhesive tape at intervals of 1 to 2 metres, and remains in the borehole after grouting. Only where the anchor is short and the borehole narrow, is the pipe inserted independently. After the borehole has been filled under low pressure, the pipe is extracted and the anchor is inserted. When the grouted section is sealed off with a collar or bag (or sometimes only a concrete sealing plug at the mouth of the borehole), air must be allowed to escape from this section through another, smaller diameter pipe leading from] the top end of the section to the borehole mouth. Outflow of grout from this pipe is a reliable sign that the section is full. This breathing pipe must then be closed before the pressure is further increased. If there is to be separate grouting of the sealing bag, the tendon

207

section, or the insulating-wrapped prestressed tendon, separate grout pipes must be attached to the anchors. In cased boreholes, the grouting of the entire borehole including the fixing section of the anchor is sometimes carried out through the casing; the casing is fitted at the top with a removable grouting head, which has a flange to take the hose leading from the pump. This method, however, is not always reliable, particularly if the anchor root is below the ground water level. Various methods for grouting long anchor roots with cement mix are discussed in Section 13.2.2. The safest method at present appears to be that in which collared tubes are used (see Fig. 13-40); these have a diameter of at least 36 mm, and are provided with lateral holes covered by rubber collars. Into this tube a steel grout pipe is inserted and sealed at the required position with a double leather seal. When grouting of the section delimited by the seal is complete and the collared tube has been flushed with water, the steel pipe with the seals is moved back to the next lateral hole, or completely removed. The collared tube remains accessible all the time. This arrangement provides, for a sequential grouting of the anchor along the entire length of the borehole, as well as allowing for further grouting at a later stage (re-grouting). The grouting pressure applied depends on the purpose of the grouting, the quality of the ground, and the depth of the grouted section below ground level. The grouting pressures for anchors usually need not exceed 2 MPa. The first filling of the borehole with cement mix prior to, or immediately after, insertion of the tendon is carried out under low pressure, or in simple cases without any pressure, by simply pouring the liquid cement mortar into the borehole. When the mix has hardened (usually within 24 hours), further grouting of the root section alone is carried out under a higher pressure. As this section of the borehole is usually closed off with a sealing bag or collar, and is situated deep below the ground surface, high grouting pressure may be applied. This secondary grouting through a collared grouting tube, may be repeated several times if need be, until the required load-bearing capacity of the anchor is attained. The grouting pressure required for anticorrosive protection of the tendon is relatively low. In order to obtain a complete filling in horizontal anchor boreholes, it is recommended that such holes be drilled with a gentle inclination (3 — 5°) away from the horizontal. Some specialized companies (such as Soletanche) maintain that there is a direct relationship between the load-bearing capacity of an anchor and the grouting pressure applied for the root fixing. It has been demonstrated in tests that the load-bearing capacity of anchors increases with increasing grouting pressure (Fig. 15-9). Similar experiments were carried out by the British company ATC Ltd. They arrived at the conclusion that when the grouting pressure exceeded 4 MPa, further increments in the load-bearing capacity were insignificant.

208 0/f

Fig. 15-9. Increase in strength of anchor fixing in relation to the grouting pressure applied [102] 1 — medium Brüssel sands, 2 — marly limestone, 3 — marls, 4 — Seine fluvial deposits, 5 — clayey gravels and sands, 6 — soft Cretaceous sediments, 7 — hard limestone

-7 ü.35

^0.25\

ψ6

•S 0.75 ^ 0-1 0.05

2

7

§.0.2

1\k

/J

3

lUnt

•

*ηκ

/m 7 injection

2 3 pressure

4 [MPa]

Higher grouting pressures are expedient in highly fissured rock, which is strengthened by the grouting, and in soil, in which the higher pressure induces a radial stress in the surroundings of the borehole, thus increasing the shear resistance at the root surface. Grouting pressures applied to soils, however, should not exceed values appropriate for the height or weight of the overburden. Littlejohn [120] recommends an admissible pressure of 0.023 N/mm 2 per 1 m of overburden height above the anchor root.

Chapter 16 FIXING OF A N C H O R S TO T H E A N C H O R E D

STRUCTURE

The method of fixing an anchor to a structure depends on the size of the structure and the anchor. Bar anchors are generally secured at the outer end by means of a nut (Fig. 16-1), while cable anchors are provided with heads clamped to the anchor ends, normally assembled after completion of the prestressing (Fig. 16-2), but occasionally beforehand (Fig. 16-3). The fixing 2

/ \

1~ T $ 3 L ^

^=±ΙΜΛ. «.--3 • Γ* Γ *

° 0

Fig. 16-1. Four threaded bars with nuts and washers bearing on a plate; the immediate bearing surfaces of the plate are angled to suit the diverging angle of the bars (system Macalloy)

1* ■ •»I

-J

**

r* '

\o" I. o

Fig. 16-2. Fixing of an anchor to a locking head after prestressing {Freyssinet system) I — sleeve, 2 — locking cone after depression

of cable anchors by concreting the ends of the cables into strong anchoring heads (Fig. 16-4) is less frequently practiced now on account of the lengthy preparation involved, although the method has many advantages to recommend it. Factory production of these heads is simple, and they are cheaper than locking heads, particularly where large prestressing forces are intended; more important than this is that these heads provide the best anticorrosive protection (see Chapter 18). A special type of end piece is the BBR V head, in which the cables are fixed by the forging of knobs at the wire ends (Fig. 16-5).

210

y

'Λ

, anchor plate /de -aerating hole vring nut

.grout pipe

B-B

A-A

Fig. 16-3. Polensky & Zöllner anchor head system (PZ) a) — before prestressing, b)— after prestressing and tightening of the nut; 1 —fixingsleeve, 2 — conically expanded end of the anchor head

46.3

Fig. 16-4. Fixed anchor head for 4 MN anchor 1 — cast steel bucket, 2 — wire ends embedded in concrete, 3 — spacing shims of various thicknesses for the purpose of prestressing

211

Fig. 16-5. BBR V anchor head system for 55 wires of 7 mm dia

16.1 A N C H O R S F I X E D BY NUTS

Nuts are used for securing bar anchors and those types of cable anchors that are provided with clamping or solid heads (see Figs. 16-3, 5, 21, 22 and Section 16.3). The threads which are to take the nuts are pressed on to the anchoring bars rather than cut, as this latter method reduces the mean anchor diameter. The nuts rest on steel washers which distribute the compressive stress over the surface of the structure or rock. The headplates which are placed on the concrete surface of the anchored structure are made of thick sheet (Fig. 16-6) and are square-shaped, with a hole in the centre for the anchor and perhaps further holes for the grouting or de-aerating pipes (see Fig. 16-3). Some systems employ specially shaped plates with turned saddles which make for the exact seating of nuts with shaped seating surfaces (Fig. 16-7). Of particular importance is the shape of the washers placed between the nut and the rock surface, because this shape markedly affects the function of the anchor and the maintenance of the prestressing of the anchor when used to secure underground excavations. Washers which are either plane or inclined at 30° to the rock surface are suitable only for the plane roof faces of galleries in bedded rocks. On the uneven faces of hard rocks in other types of underground excavations, the nuts or bars may be seated excentrically if such washers are used; this creates unfavourable loading of part of the bolt by a bending moment, and reduces its load-bearing capacity. In addition, the prestressing operation is more difficult and errors may occur in the application of torque spanners. Accurate centering of the bolt in the borehole and uniform distribution of the transmitted force over the rock surface can only be

212

achieved with supplementary spherical-surfaced washers, currently supplied by some specialized firms. The washer itself, bearing on the surface of the rock and usually holding in place a protective netting, is suitably formed so as to improve its function and economise on material (see Fig. 16-7).

Fig. 16-6. Dywidag distribution plate system a) — plate located on the surface of the structure, b) — plate embedded in concrete, levels with the surface of the structure, c) — hollowed saddle in the seating plate ensuring precise seating of the nut

Fig. 16-7. Washer and bolt nut supplied by Pneumatisk Transport AB, Sweden

The South African firm of Roof bolts S. A. Ltd. has developed a special washer for fixing bar anchors on to rock surfaces. This washer is slipped on to the unthreaded bolt rod, and after prestressing, the washer grips the rod strongly and reliably by a self-locking effect; this fixing thus belongs to the next category of anchor heads (locking heads). The triangular curved shape of this washer guarantees support at three points on the rock surface. A given tilt of the thickened part of the washer with respect to the anchor rod corresponds to a particular tensile force in the anchor; thus, the washer also acts as a check on the stress within the bolt (Fig. 16-8). Bolts fitted with these heads are capable of supplementing prestress, if required.

213

Fig. 16-8. Self-locking and indicating washer of Roofbolts S. A. Ltd., Johannesburg prestressing a) 0—50 kN, b) 50— 60 kN, c) 60—103 kN; over 100 kNi/), fixing of washer breaks down

16.2 L O C K I N G HEADS (WEDGE B L O C K I N G SYSTEM)

Locking heads are mainly used for fixing cable anchors composed of straight wires, or single and multiple strand anchors. The locking effect is obtained by means of wedges or truncated cones, which are pressed among the wires or cable strands and forced during the prestressing into tapered holes in steel bearing plates. Locking heads (functioning by the effect of friction) do not project on the surface of the structure. The length of the cable can easily be adapted according to the dimensions of the structure. Manipulation of this type of head is simple, no time is needed for the concrete to harden, and these heads do not require the special attention demanded by other systems (e.g. protection of the threaded parts from damage during transport and fitting). For the anchoring of cables composed of straight wires, heads consisting of a sleeve and cone are used; the cone is pressed among the wires and into the sleeve by hydraulic force when the required prestressing is applied (Fig. 16-9, see also Fig. 12-4). This system does not permit any post-fixing adjustment of the anchor prestressing, a problem which has been solved, for example, by the firm of Polensky & Zöllner (GFR) who secure the wires in a reversed anchoring head (see Fig. 16-3). The anchoring cone is tapered outwards, continuing as a screw to which the stressing equipment is attached; the screw also holds the nut which secures the anchoring head in its final

214

^?jfiSi

Fig. 16-9. Anchor head for 1 MN anchors Horel system in service on a rock slope

A Fig. 16-10. Losinger anchor head system (VSL) 1 — cone, 2 — grout intake, 3 — hole to facilitate assembly, 4 — elongation -^ 100 mm, 5 — fixing sleeve, 6 — securing ring with recesses for spanner

position after prestressing. The prestressing of the anchor can be altered if desired after the fixing has been made. The anchoring head of the Losiwer firm (Fig. 16-10) represents another arrangement with the same facility. On prestressing, the cone is automatically drawn into the sleeve among the wires; the prestressing equipment is attached

215

by a securing nut to the outer threaded surface of the sleeve. When the required anchor prestressing is reached, the ring-nut is tightened with a spanner against a strong washer on the anchored structure and the prestressing equipment is disconnected from the anchor head. In neither of the arrangements mentioned above are there prestressing losses as a result of slipping, as occurs in locking heads when the cable end is fixed by the action of the cones; the prestressing can also be easily adjusted at a later stage. In spite of the fact that the fixing equipment is more complicated and requires careful maintenance, both of these methods are in common use. The same type of head is sometimes used for the fixing of anchors composed of strands (Fig. 16-11). In order to ensure reliable fixing of each of the strands of the cable, various modifications have been devised to prevent the slipping of individual strands. 0W7

t

/

Fig. 16-11. Monogroup anchor head

Fig. 16-12. VUIS segmented anchor head, loading capacity 1 MN 1 — segments, 2 — locked strand space between segments

«^^^ ^^^^M

y^

T

1βκ1.5

3

f^^f^

->■

\—ΛΤ

"

,

\

4

Segmented heads of the VUIS type are used in Czechoslovakia for the fixing of multiple rope anchors. Using these, load-bearing capacities of up to 1 MN may be obtained for seven strand anchors ( 7 x 7 wires diameter 6 mm; or 7 x 19 wires diameter 4.5 mm (Fig. 16-12)). The PSC head, used in Great Britain for fixing cables, follows the Freyssinet system and consists of a conical ring into which a bevelled cone fits. On the inner surface of the ring and on the surface of the cone, grooves are formed so as to interlock with the anchoring rope surface when the rope is clamped in the grooves (Fig. 16-13).

216

The CCL system (Fig. 16-14) can be used for various types of multi-strand cable, but cables composed of seven strands, each 12.7 mm in diameter, are recommended as the most suitable type, although cables composed of a larger number of strands can also be accommodated. A special feature of the system is that each rope is prestressed and fixed separately, using light-weight equipment in which the end of the rope passes through the axis of the prestressing jack cylinder. The system developed at the VUIS, Bratislava (Czechoslovakia), in which the cable is prestressed by taking pairs of ropes, has a similar design (Fig. 16-15). Another proven type of head consists of a plate against which cylindrical components (collars) attached to each of the individual ropes abut (Fig. 16-16). 108+153 mm

Fig. 16-13. PSC — FreyssiMonogroup cable head system. The head dimensions depend on the number of strands fixed. The head is designed for cables with a maximum loadbearing capacity of 8 MN

cables 7/12.7mm 7/15,2mm 7/17.8 mm

Fig. 16-14. CCL — Multiforce cable head system

O Fig. 16-15. Anchor head in which strands are stressed in pairs {VUIS system)

217 Fig. 16-16. Fixing of the cable

Fig. 16-17. Stress Block head fixing system for 21 strands

Fig. 16-18. KA head fixing system

In these systems the individual strands are prestressed one by one with light prestressing guns, or all together with a set of several such guns (Chapter 17). Individual prestressing of the cables is also a feature of the Stress Block system (Fig. 16-17), which differs from the latter systems only with respect to the arrangement of the cable ropes and the shape of the head. Ropes 12,7 mm in diameter are assembled in triads into bundles of up to 45 ropes (5.1 MN) with a rectangular cross-section overall. Parallelism of all the cable ropes is ensured by using spacing grids, which also make for uniform envelopment of all the ropes with grout. In all the types of head described in the foregoing, the anchorage was fastened to the head by the locking effect of various shaped wedges, cones, or segments, inserted so as to press the wires or strands of the anchor against

218

an outer sleeve member. Recently, however, heads have been introduced in which wires and steel bars are locked by compression in a system of steel shims between two bolted plates (Fig. 16-18). Thus in the KA system, a high fixing strength is obtained by using wires of oval cross-section for the cables. These wires have a ragged surface which corresponds with a similar ragged surface on the shims and locking plates. The manufacture of BBB heads, designed to take an even number of strands (2 to 12) of 12.5 mm diameter, is very simple. These heads are made of cast iron and have bulb-shaped cavities; the strands are inserted in pairs and are pressed against the cavity walls by a wedge placed between them. Heads for cables composed of 2 or 4 ropes are concreted in advance at the borehole mouth (Fig. 16-19a).

Fig. 16-19. BBB anchor head system a)—head embedded in concrete for 2 or 4 strands, b) — star-shaped head for 12 strands

The head for 12 ropes is star-shaped and rests on a mild steel bearing plate cast into the concrete (Fig. 16-19b). The wedges or conical segments are designed and inserted into the stressing equipment in such a way as not to obstruct release of prestressing tension after completion of the prestressing; this is essential, particularly when soil and rock anchors are prestressed (see Section 17-6). The heads are padded only in those cases in which anchors have to be later re-stressed in order to compensate for losses of stress caused by creep of the ground.

219 16.3 IMMOVABLE ANCHORING HEADS

Immovable anchoring heads were used in the early days of the application of rock anchorage, when large cable anchors were fixed to the surfaces of anchored structures; however, with modifications and simplification of assembly, these heads are still useful in many present-day types of anchorage. Immovable heads are made either of prestressed concrete, or of steel cast in a shell-like form, and are filled with white metal or concrete after having been mounted on the external end of a bundle of wires. Cable ends such as those of cableways and suspension bridges, can also be fitted with these anchor heads. The heads are conical inside. The fixing of wires is achieved by cohesion and the transverse pressures created by the pulling-in of the cone of wires and white metal into the cast steel head. Alloys with a smelting point of not more than 330 °C are used for the cast around the wires, otherwise the wires may be damaged by overheating. Steel buckets filled with grout are often used to make strong anchor heads, and good results have been obtained with these at Czechoslovak dam sites for the surface fixing of steel ropes, prestressed to 4 MN (see Fig. 16-4). The buckets are 50 cm deep, have a seating area of 45 cm diameter, and an 11 cm diameter opening for the cables. A reinforced concrete anchoring head may also be created by embedding the spread end of a cable consisting of straight wires in concrete (Fig. 16-20). The transverse tensile stresses in the head and the shear stresses brought about by the pull of the anchor tendon are taken up by helical (and also sometimes radial) reinforcements. The manufacture of these reinforced concrete anchor heads is a labourious process, and heads of this type are only suitable for the largest cable anchors, with carrying capacities of 7 - 1 2 MN (see Section 24, the Cheurfas Dam). All types of immovable head must be supported on suitable headplates in position before the anchor is prestressed, although some types of fixed anchor head are provided with a thread and nut for that purpose. In the SEEE system (Societe d'Etudes et d'Equipements d'Enterprises), special heads are used which are fixed by being pressed on to the outer end of the anchor cable (Fig. 16-21). The rope ends are inserted into thick-walled steel tubes of certain specified properties, and the tubes are then pressed to a smaller diameter. Thus, for example, a tube 500 mm long and 108 mm in diameter, after being pressed, forms an anchor head 620 mm long and 94 mm in diameter. In some cases a helical coil is inserted around the rope in the tube in order to increase the strength of the fixing. When the tube has been pressed on to the rope, a fine thread is cut or pressed on its surface. Following prestressing of the anchor a nut is screwed on to this thread and tightened against the load distribution plate on the front of the structure. SEEE pre-

220 Fig. 16-20. Reinforced concrete anchor head for cable of straight wires 1 — head, 2 — load-distributing steel plate, 3 — steel washers, 4 — reinforced concrete props W2 Φ5ΓΠΓΠ

MM

WKKKeKHmKm Fig. 16-21. SEEE system heads of different lengths

view of the stressed rope end prepared in the structure 1

Fig. 16-22. Baudin— Chäteaimeuf fixing system / — rope, 2 — resting plate, 3 — anchor head, 4 — stressing head, 5 — helix reinforcement, 6 — casing, 7 — cast-in metal, 8 — compression of ropes by jaws

stressing units are designed to withstand forces of 1.12 and 2.75 MN. In the first case they are constructed of a single rope with 61 regular lay wires of 4.1 mm diameter, and in the second case the unit consists of 19 ropes, each of7 wires 3.6 mm in diameter. The anchoring heads of this system can be prepared entirely in the workshop; their assembly for prestressing, and the initial and subsequent (adjustment) prestressing operations are simple. The Baudin —Chäteauneuf system (Fig. 16-22) was also developed in France. Here a special type of head is used for fixing ropes. To form the head, the rope is held at two points in the jaws. When these are closed somewhat5

221

the wires of the rope between the two holding positions are loosened and a bulge is formed; the bulge is then covered with a thin-walled protecting tube and the wires are cast inside. Prestressed units of this system have a loading capacity of 5 MN or more, and are supplied to the site, complete, in coils. The BBR V heads are much used on structures made of prestressed concrete. They are designed for cables composed of wires and are mounted on the anchors during manufacture in the workshop. The ends of the wires are slipped through holes drilled in the anchor head, and cold forged in specially designed equipment. When long anchors are assembled, the heads are inserted into the expanded mouth of the fixing hole to a depth such as to obviate padding of the head with thick plates after prestressing. If the depth of the head has been correctly calculated, the head will be located at the surface of the structure after prestressing. Safe seating of the anchor is achieved with the aid of a ring screwed on to the anchor head. The new position of the anchor head after additional prestressing is also set by turning this ring. For shorter anchors, the heads are padded after prestressing with annular washers split in halves (Fig. 16-23). a)

b)

Fig. 16-23. BBRVanchor head a) — for long tendons with large stressing elongations. (In the non-stressed condition of the tendon, the anchor head is located inside the trumpet), b) — for short tendons with a limited elongation. After prestressing the anchor is locked with a pair of shims

16.4 SADDLE ANCHOR HEADS

In some cases an upper movable anchor head is replaced by a concrete metal-clad saddle (Fig. 16-24). A cable is led over the centre of this saddle, and both ends are fixed into the rock. When the cable is prestressed, a jack is placed under the centre of the saddle, and is replaced on completion of the

222

prestressing by reinforced concrete and steel blocks. This method appears to be simple enough, especially as the number of cable ends that have to be manipulated at the head is halved. There is, however, a disadvantage in that the boreholes into which the cable ends are fixed must be located close together, and this means that the resistance to extraction of the anchor from the rock does not greatly exceed the extraction resistance of a cable fixed in one borehole. Because of this, the use of saddle anchors necessitates doubling of the borehole length, thus making the method uneconomic. In concrete structures, in which anchoring holes are formed by embedded pipes, this disadvantage may be avoided by arranging the embedded pipes so that they diverge away from the opening of the common hole in the footing of the anchored structure, which leads to the fixing head of the anchor. Saddle anchors can be used to advantage where the cables are laid on each side of the anchored structure (see Fig. 24-2).

Chapter 17 P R E S T R E S S I N G AND T E S T I N G OF A N C H O R S

The purpose of prestressing an anchor is to create an elastic tension in the free section of the steel anchor tendon with the aid of suitable stressing equipment; in this way, the tendon section exerts a predetermined force on the anchored structure. The prestressing of an anchor provides a test of the anchor at the same time. It confirms the suitability of the anchor type to some extent, it suggests what future behaviour of the anchor can be expected and it indicates errors in the design and installation of the anchor. Changes in the characteristics of the ground and deviations from the anticipated installation conditions can be the cause of substantial differences between the load-bearing capacities of anchors at one and the same construction site. Hence, it is important to subject each anchor at the site to a test. The methods of prestressing, testing, and checking anchors are now fixed by Standards and Codes in many countries [233 — 241]. The Standards and recommendations differ in detail, but the basic procedures must take account of the characteristics of the materials used and the safety demands of the anchored structure, and are therefore the same everywhere.

17.1 STRESSING FORCES

Anchors may be stressed to the production load or the testing load. The production load of an anchor is given by the working (admissible) force, Pw9 calculated according to the static analysis; the anchor must be able to sustain this force throughout its entire service life. The production stressing of an anchor usually corresponds to this force. The working force must be extended by some safety margin before the anchor's ultimate state is reached, as determined by the point of failure of one of its main components (breaking strength failure), or the exceeding of the admissible deformation (e.g. yielding failure of the tendon steel). The safety margin is determined from the results of basic anchor tests, or it is laid down by a standard code of practice drawn up in the country concerned. A range of safety factor values, compiled from accessible Standards and recommendations, is listed in Table 17-1:

224 TABLE 17-1 Safety factors for establishing Pw With regard to the ultimate strength of anchor steel

1.65—2.00

With regard to the yield strength (and/or elastic limit) of steel

1.33—1.65

With regard to the ultimate strength of the anchor root in rock or soil

1.60—1.70

In some countries a distinction is made between a temporary and a permanent anchor, in determining Pw. The safety factor for temporary anchors is usually lower by one or two tenths than that for permanent anchors. The testing load is a short-term loading applied to the anchor in order to test the integrity of the whole installation, check the safety factor chosen, and make sure that the anchor has the capacity permanently, to transmit the working force, Pw; alternatively, the ultimate strength may be measured by the test carried out in a given type of ground. The testing load generally reaches higher values than the production load. The maximum testing force, Ptt max for production anchors is established as a function of the working force, P w , or is given by the limit states of the tendon steel-the tensile or yield strength (and/or elastic limit). Values of testing forces are introduced in the next Section. A.max i s usually the largest stress experienced by the anchor during its service life, with the exception of circumstances in which the anchored structure or ground is overloaded by forces which had not been anticipated in the static analysis. Besides checking the installation, the higher testing load to some extent makes good losses of prestressing which appear in the anchor tendon as a result of compression of the ground or anchored structure the fixing procedure of the anchoring head and headplates, relaxation of the tendon, etc. In special tests on anchors, P f , m a x is determined by the ultimate load of the anchor, until failure occurs.

17.2 STRESSING^OF P R O D U C T I O N ANCHORS—ACCEPTANCE TESTS

The stressing of all production anchors is carried out in the form of an acceptance test, in which the anchor is subjected to a test load greater than the working force, Pw, for a predetermined time. The purpose of the short-term loading of an anchor by a larger force is to obtain measurable safety coefficients relating to the designed production

225

load, Pw, or to discover defects in the design or installation of the anchor in good time. In the test, the displacement of the anchoring head is measured as the tensile force is increased, and the displacement continues to be monitored throughout the period of the test. The force-displacement diagram gives information about the operating characteristics and parameters of the anchor (e.g. the free tendon length). The displacement-time diagram gives an indication of the integrity of the anchor fixing in the ground. The maximum value of the testing force for production anchors is generally laid down by different national Standards, or it is chosen according to recommended values. For permanent anchors, the testing force is usually higher than that applied to temporary anchors. Values of Pu max , compiled from available Standards and other sources, are given in Table 17-11. TABLE 17-11 Maximum testing force, Ptt max, for production anchors Permanent anchors

A.max = 1-20—l.50Pw = 0.70—0.85PS = 0.90—0.95Py

Temporary anchors

Λ,ιηκ = 1.15—1.25Λ, = 0.70—0.85PS = 0.90—IMP,

P» is the ultimate tensile force for tendon steel, Py is the yield-point force (or elastic limit force) for tendon steel. (The yield-point is usually defined as 87 per cent, of the tensile strength, the elastic limit as 83.5 per cent, of the tensile strength)

Simple acceptance test The current stressing procedure follows that of the simple acceptance test. The procedure is as follows: The anchor is stressed by an initial force P0 = 0APw to 0.2PW (0.1 P y ), and the initial reading at the anchoring head is registered. The loading of the anchor is progressively increased to Pu max = \A5PW to 1.5PW, the displacement of the anchoring head is registered, and subsequent further displacement of the anchoring head under this load is monitored for a minimum test period. The test period is usually 5 to 15 minutes, according to the standard adhered to, and sometimes depending on the type of rock or soil in which the anchor is fixed. For strong rocks 5 minutes are sufficient, whereas for cohesive soils 15 minutes should be considered a minimum. Usually it is

226

demanded that stabilization of the anchor (no further change in the displacement) has occured by the end of the test period (see more detailed test), or that a predetermined total allowable displacement has not been exceeded ( 1 - 2 mm). If this condition is fulfilled, the anchor load is decreased to P 0 , and the drop in the displacement at the anchoring head is recorded, thus giving the elastic and permanent components of the total displacement. Then the anchor is loaded by the working force Pw9 increased by the assumed losses caused by friction and relaxation (usually 0.1PJ, and the anchoring head fixed. The results of the test are presented in a report, which contains all data concerning the dimensions and installation details of the anchor, and the values obtained in the test. The report includes the load-displacement diagram, plotted from the measured values (Fig. 17-1). The total displacement is divided into permanent and elastic components. If the measured elastic displacement lies between the points Bx and B2, then the free length of the anchor tendon corresponds to the designed length and the transmission of forces to the anchor root corresponds to the assumptions made in the design. The two points, Bx and B2, represent the calculated elastic deformation of the tendon free length, reduced by 20 per cent (point BJ, and increased by a half of the root length (point B2). For anchors in which the tendon is fixed to a steel base at the extreme end of the root and is insulated up to p

t-ma*

l0Qd

P

permanent displacement line

boundary fine 1

elastic displacement tine total displacement line boundary line 2

B2

Fig. 17-1. Load-displacement diagram obtained during a simple acceptance test of an anchor

227

this fixing point, the points Βλ and B2 are given by the free length of the tendon multiplied by-4he coefficients 0.9 and 1.1, respectively. The elongation of the steel tendon under a given load is calculated from the equation P.L where P = tensile force acting on the anchor, L = the free length of the tendon, multiplied by the coefficient for Bx or B2, E — the modulus of elasticity for tendon steel, A = the area of cross-section of the tendon steel. Detailed acceptance test Most Standards stipulate detailed tests on a minimum number of production anchors. Usually this concerns the first (3 —1.0) anchors installed at the site, and thereafter one or more anchors from each group of 10 to 25 subsequently installed, or a percentage (5 per cent, on average) of the total number of production anchors at the site. In this acceptance test the loading starts with the initial force P0, and is increased gradually to the maximum testing force. Usually there are 4 to 5 loading steps, for example 0.4PW, 0.8PW, 1.0PW, 1.2PW, (1.4 or l.5Pw). At each step the anchor is relieved as far as P0, and when the residual deformation has been measured, the loading is increased to the next step. At each loading step the force is maintained and the displacement at the anchor head is measured until stability is observed. In the last loading step the observation period is usually longer, being 1 to 24 hours according to the type of rock or soil in which the anchor root is fixed. The displacement must be recorded äs having stabilized, or a maximum overall rate of displacement must not be exceeded, as in the previous test. It is recommended that the displacement increment be measured with a dial gauge at progressively 1 3 1 1 increasing time intervals, e.g. 1 — , 1 — , 2 — , 3 — , 5, 7, 10, 14, 20, 28, 40, 56, 80, 112, 160 minutes, etc. Stabilization of the displacement can be considered as having occurred when the displacement increments measured at the anchor head remain unchanged over three successive time intervals, or if they change by no more than 0.1 —0.2 mm. Other details of the procedure and the graphic expression of the results are similar to those of the simple test. Besides the load-displacement diagram (Fig. 17-2), the displacement-time diagram is often drawn as well. If the time values are plotted on a logarithmic scale, a displacement movement that is

228

approaching stability will be represented by a nearly straight line. From the time-displacement diagram the approximate prestressing losses R, due to friction during the stressing of the anchor can be determined also (see Fig. 17-4). a

j A^a1p*

^ g *,*—

^gS^X

\< & -^ §

1

^ ^ ~^ ^ ^ .U ^

°-8Pw Pw

<x £x^ ^vvv ^ *WCsX ^vx v \ ^§^ΝΧΚΝ

permanent displacement line X . X

\NsJ^s x χΟ\

-+Ξ

£

? —

ί

1

! '

1-2PW 1MPW load

Cvfcv

CM

Λ

MPw

Γ

-+^

sξ

\J Vl *3

"I1 ^'

s

X

\

§

V

x 1 boundary line 1

elastic displacement line Nl \ ^xV~. \ X total displacement ^ X i line X>J Xl

C \^ NX

N

boundary line 2

Fig. 17-2. Evaluation of a detailed acceptance test of an anchor according to K. Klein [235]

17.3

SPECIAL A N C H O R TESTS

Special tests are carried out on those anchors which are not used as production anchors, but are perhaps prototypes for researching new anchoring systems, or test anchors used to verify the suitability of a known type of anchorage for a particular site. These tests are called basic tests in the former case, and suitability tests in the latter case. In both cases the behaviour of the anchorage specified the design is examined, particularly the safety margin afforded against failure; the tests should also reveal any defects which might influence the load-bearing capacity of the anchors during their long-term service. Basic tests The technical suitability of a new type of anchor and the method of construction of anchors for temporary and permanent use, must be verified by conducting a basic test in the type of ground in which the anchor will be used (strong rock, soft rock, loose soil, cohesive soil). It must be demonstrated that the anchor satisfies the expectations of the design, or the stipulations of a given Standard.

229

In every type of rock or soil for which the anchor has been designed, at least three anchors must be tested. The anchors are loaded in stages as in the detailed acceptance test, the stages being established as functions of the yieldpoint force of the tendon, and increasing, for example, in steps of 0A5Py. A constant force is maintained at each loading stage and the displacement at the anchoring head is measured until it becomes stable. All forces are measured with special instruments (see Chapter 19) and displacements with dial gauges of 0.01 mm accuracy. Displacements are registered for a period sufficiently long to take account of creep of the ground. For example, the German Standard recommends that displacements occurring under the higher loading stages be measured for at least 2 hours where the anchorage is made in coarse-grained soils, 24 hours in fine-grained (cohesive) soils, and in any case until the displacement increments are less than 0.2 mm. From the load-displacement graph the force, Pc9 under which the measured deformation has attained 2 mm (Fig. 17-3), can be found. The force Pc is also used as a criterion for finding the safety factor, as will be explained later.

aww u

o.dPw 1.2PW 1.5PW

fen sile

load

Fig. 17-3. Determination of the limit anchoring force, Pc [120]

The load in a basic test is increased until failure of the anchor occurs, and this should take place in the root, not in the tendon. Failure occurs when the displacement increases with time under a constant loading force. On the basic test graph of Fig. 17-4, the load-bearing capacity of the anchor root was exceeded under a load of 0.94P y , and the free tendon length behaved according to expectation. When the test is completed, the anchor is dug out and examined. Particular attention is given to the root, its shape and dimensions, the quality of the hardened grout, the contact established between the tendon and the grout, the position of the tendon in the grout body, the covering of the steel parts by the grout, the appearance, number, and mutual spacing of cracks in the grout body, and the effectiveness of the anticorrosive protection of the root and tendon, particularly in the case of permanent anchors.

230

-^displacement a)

elastic displacement - * M

-^permanent displacement

Fig. 17-4. Evaluation of the basic test of an anchor, according to DIN [236] a) — stress/strain diagram, b) — curves of elastic and permanent displacements. At a loading of 0.75Py, stabilization of the displacement was not registered. R designates the loss of tensile force involved in overcoming friction during stressing of the anchor

The complete results of the test are contained in a report which also gives the characteristics of the ground, the anchor parameters, the test method, and a discussion of the test values in relation to what has been found in the examination of the excavated anchor. The construction, testing, and subsequent excavation of the anchors must be supervised by a recognized professional body, which also classifies the ground at the test site. According to the German Standard, the admissible working load, Pw9 of a given anchor type is taken as the smallest of the values P y /1.75, Py/1.75, PJ1.50, obtained from the basic test. Pw is taken as 0.50Pf for permanent anchors, and as 0.67Pf for temporary anchors, according to the French Standard. Py is the force at the guaranteed yield strength of the anchor tendon, Pf is the load at the failure limit of the root as found, in the test, and Pc is the test force that produces a displacement of 2 mm of the anchoring head. Suitability tests The purpose of these tests, as has been stated earlier, is to check the suitability of an anchor type for use at a given site by subjecting it to harsh or little-known conditions. The suitability tests are performed on site before construction work is started. Usually three test anchors are positioned where the least favourable geological conditions are assumed to exist, and they are subjected to loading and release cycles with simultaneous recording of the displacement at the head, as in the basic test. From these results, the elastic

231

and plastic deformations are obtained, and then the actual free length of the tendon can be calculated. The plastic deformation approximately corresponds to the displacement of the root [154]. The particular linear relationship between the displacement and loading on an anchor characterizes the loadbearing capacity of that anchor. The maximum load test is usually fixed at 1.5 or 1.6 times the design working force, Pw, for temporary and/or permanent anchors; sometimes the anchors are loaded until the ultimate load is exceeded, but then they are not dug out afterwards. A detailed certificate is drawn up from the test results. On a site where, according to the design, the anchors are to be placed less 1 m apart in a line, the possibility of their mutual interference must be investigated. In this case, test anchors are installed as the design specifies, and are loaded simultaneously while their behaviour is observed.

17.4 TESTS ON ROCK BOLTS

g

17A.1 Destructive tests Pulling tests are usually carried out on rock bolts as an aid for the selection of bolts, most suitable fixing materials, and installation methods. Pulling tests are based on the same principle as the basic testing of long anchors, the applied load obviously being smaller. The bolts are loaded in stages over a short time interval until failure occurs, and the changes of displacement of the bolt head are measured simultaneously. The result is plotted in a loaddisplacement diagram, which represents the behaviour of the bolt in the same way as that of a prestressed long anchor in rock. The test is destructive and therefore cannot be carried out on bolts that form a part of the actual rock reinforcement or support system. A method for determining the strength of rockbolts by applying a pulling test was given, for example, by the International Society for Rock Mechanics [95]. The main principles are: 1. At least five tests are required to evaluate an anchor in a given set of rock and installation conditions. 2. The loading equipment is assembled, taking care to ensure that the direction of the pull is coaxial with the bolt, that the equipment sits firmly on the rock, and that no part of the bolt or hardened grout will interfere with the application or measurement of the load during the test. 3. An initial arbitrary load not greater than 5 kN is applied to take up any slack in the equipment. The displacement measuring equipment is assembled and checked.

232

4. The anchor is tested by increasing the load until a total displacement of more than 40 mm has been recorded, or the bolt yields or fractures first. 5. The load and displacement are recorded at increments of approximately 5 kN load, or 5 mm displacement, whichever comes first at each step. The rate of loading should be within the range 5 — 10 kN/min. Readings are taken only after both the load and displacement have become stable. The times required for stabilization to take place should be recorded. 6. The test data are plotted graphically, as shown in Fig. 17-5. The anchor strength, defined as the maximum load reached in the test without the bolt yielding or failing, is recorded in this graph. If the bolt yields or fails, the load "
3 2

i\

}^

** \

j

A

\\\iI /

( Γ

/ /

\

Ϊ<100

*ί \ \

uv\ LL

20

10 20 30 Ϊ0 50 bolt head displacement [mm] Fig. 17-5. Example of graphic plot of bolt test results [95] 1 — elastic deformation line, 2 — yield-point load, 3 — ultimate load

^ΜΏ ΤΓΓΠΤΤΤΜ

120

■^>

ξ>80

/

-Η

r*i————

^ 60

} II / LLrj ι

Ι4-πΤΓ

20

0.5 10 15 2.0 bonded length [m]

I

2.5

Fig. 17-6. Graph showing the influence of bond length and curing time on the strength of bolts. (The results shown are hypothetical) [95] 1 — five-day cure, 2 — one-day cure

7. The elastic elongation s of the bolt at a given applied load, P is given by P .L where L is the tensioned, ungrouted length of the bolt, plus one third of the grouted length, plus the length of the extension bar if used, A is the cross-sectional area of the bolt, E is the modulus of elasticity of the bolt steel. A straight line (1) is constructed to connect the origin of the load-displacement graph with the point on the graph representing the load, P, and corresponding displacement, s. Straight lines (2) and (3) are drawn at the theoretical

233

yield-point and ultimate loads of the bolt. Comparison of the actual test curve with these three lines, leads to an understanding of the behaviour of the bolt and its anchorage. 8. For the evaluation of grouted anchors, the results of several tests should be abstracted and presented graphically to show the influence of the grout setting time and the length of the bonded section on the anchor strength (e.g. Fig. 17-6). 9. The report of the test should include data sheets and graphs, together with full details of: a) the rock in which the anchors were tested, b) the anchors and associated equipment, c) the drillholes, including the length, diameter, method of drilling, straightness, state of cleanness and dryness, orientation, d) the method and time taken for installation, e) the method and time taken for the testing, f) the nature of any failure, and other observations pertinent to the test results. 17.4.2 Non-destructive tests The first non-destructive method for in situ testing of rock bolts, especially grouted bolts, was p r esented by the Geodynamic AB Stockholm. A compact, easy-to-use field instrument, based on electronic technique (Fig. 17-7), is

Fig. 17-7. Non-destructive test of Geodynamic AB applied to a rock bolt

234

capable to detect if a bolt is cut, if grouting has not enveloped the entire length of the bolt, if the bolt is too short, if the contact between the bolt and grouting or between the rock and grouting is defective, and also if the contact between the expander and the rock is insufficient for the mechanical fixing of the bolt. The instrument has a head which contains a transducer — senzor of piezo-electric crystals. The head is held against the exposed bolt end and the transducer transfers elastic waves to the bolt. These waves propagate down the bolt at different rates (depending on the bolt length, grouting and contact conditions) and are reflected back to the sensor. By processing the detailed signal informations received, both bolt length and contact conditions can be determined. 17.5 REQUIREMENTS OF THE EQUIPMENT USED FOR THE PRESTRESSING AND TESTING OF ANCHORS

The equipment used for prestressing is practically the same as that used for testing anchors. It consists of a hydraulic set, stressing head, anchoring head, and a part of the tendon at the near end of the anchor. Sometimes it also contains a load sensor for precise measurement of the tensile force, and a dial gauge for precise measurement of the tendon displacement (Fig. 17-8). The hydraulic set comprises a pump, a jack, a connecting hose, and pressure gauges of appropriate range and accuracy (1 — 2\ per cent.). The stressing equipment must be capable of creating a tensile force in the anchor tendon, and must maintain this force at a constant value. The value of the stressing force acting on the anchor tendon must be measured by sufficiently accurate instruments checked by a recognized authority at regular intervals. It suffices to measure this force with two calibrated pressure gauges of the required accuracy, connected to the hydraulic system between the pump and jack. Measurement of the stressing force with a dynamometer mounted on the anchor tendon is more reliable and more precise. Approximate measurements of the tendon displacement can be obtained with the scale fixed to the jack piston. Precise values are obtained with a dial gauge, which should be mounted on a supporting structure independent of the prestressed object, so that movements caused by prestressing are reliably registered on the gauge. To establish the relationship between stressing force and displacement, it suffices to measure the tendon displacement to the nearest 0.1 mm; to obtain the displacement-time relationship at a constant stressing force, the measurements of the tendon displacement must be accurate to 0.01 mm, because the tendon displacement measured at the anchoring head may be small, yet continue to change over a long period on account of creep of the anchor tendon, anchor root, or the rock or soil.

235

The measuring instruments must be kept in good working order and must be regularly calibrated to ensure reliability of the readings. It is advisable [120] to have the measuring instruments calibrated before each stressing operation and to check them on the spot against control instruments at monthly intervals, or after every thirtieth production anchor installation. An independent calibration of the stressing equipment should take place every three months. Even with the greatest possible accuracy and care, errors in the measurements nevertheless occur on account of inaccuracy in the pressure gauges, internal friction in the jacks, inaccuracy in the displacement gauges, and particularly on account of the manufacturing tolerances of the anchor tendon. The difference between the actual (measured) and calculated values of the stressing force may be up to 15 per cent, for anchors, and generally averages 5 per cent, in current installations [120]. In the stressing and testing of anchors, considerable forces are employed. Consequently, the area immediately behind the stressing equipment must be

Fig. 17-8. Equipment for stressing and testing anchors (A — Brückner Grundbau Co.,)

236

I

Precision _/ manometer

Smooth duct

Fig. 17-8. (B, C — Losinger Ltd.) in service on test sites for the precise measurement of load and displacement

237

cleared for safety reasons; and access to this*«rea must be prohibited. Serious injury could otherwise be inflicted by sudden failure of the anchor, or collapse of the stressing equipment.

17.6 PRESTRESSING TECHNIQUES

The method of prestressing an anchor and the design of the prestressing equipment, both depend on the type of head used at the accessible end of the anchor, and the magnitude of the stressing force. It is very important that the stressing force always acts in the direction of the tendon axis, and that no bending of the tendon can therefore take place. This can be achieved by means of a suitable load distribution plate mounted on the surface of the anchored structure or rock so as to keep the anchoring head oriented coaxially with the tendon axis. 17.6.1

Prestressing of bolts

Short bar anchors (bolts) with only a minor degree of prestressing are often tensioned by simply tightening the securing nut. The torque needed to turn the nut on the bar anchor does not indicate directly the tension within the anchor; calculation of the latter depends largely on the frictional resistance of the nut on the thread and washer in the course of tightening. A. Hugon and A. Costes [94] expressed this relationship by the formula: C ^ - ^ - ( t g j 9 + 2tg^), where C = the torque (Nm), P = the tensile force (N), d = diameter of the bar, excluding the thread (m), ß = angle of the threading, φ — angle of friction of the nut on the thread and washer. With the metric thread currently used for bolts, the average value of β is 2°30\ but in the Ancrall type is 9°, and in the Pattin type, 5°. The median value of the friction angle, φ, (according to Hugon and Costes) is 14°, assuming that the thread and the nut are in good order. The graph shown in Fig. 17-9 gives the necessary turning moment, calculated from the above equation, for producing a given tension in bolts of the most frequently used diameters (20 and 24 mm, with M 20 and M 24 metric threads). Average figures for the tensile force obtained by direct measurements on bar anchors, were found to be approximately 20 per cent, lower than the calculated

238

values. This discrepancy can be removed by introducing a friction-reducing material beneath the lock-nut prior to stressing. In countries where the Whitworth thread and older units of measurement are still in use the torque-tension graph in Fig. 17-10 will be more convenient; it is published by the American Williams Co. for bolts of various diameters.

50

100

150

200

250

300 350 400 450 torque moment N/m

500

Fig. 17-9. Relationship between the torsional moment on the securing nut, and the tension in the bar, for bolt bars with M 20 and M 24 threads

Fig. 17-10. Torque-tension graph for bolts, according to the American Williams Company

In mines and tunnels bolts are often prestressed by hand, using a flat spanner with its arm extended to 80 or 100 cm. If a hand force of approximately 0.3 kN is attainable at the end of the spanner (although this varies considerably, of course, according to the position of the workman doing the tightening), the resulting torque is 235 Nm, which corresponds to a tension of 39 kN in a bar 24 mm in diameter with a metric thread. More accurate tensioning of bar anchors is achieved with a torque spanner (wrench). Torque spanners for tightening nuts are currently made from light alloys,

239 and have various measured moment ranges. Torque spanners producing moments of up to 700 Nm are most often used for prestressing bolts. They are equipped with exchangeable ratchet adaptors. For example the well known Spanish Torcometro torque spanners (Fig. 17-11) which are made in three sizes. W0wiiMBMMk:':'i:i'

Fig. 17-11. Torcometro torque wrench

The largest type, with a torque range from 275 — 740 Nm (weight 3.2 kg, length 80 cm) can also be used for tightening the bolt nuts to achieve the required prestress. The required moment (in Nm) is set on the scale of a screw gauge on the handle of the spanner; a click indicates when the moment has been reached. A more suitable instrument for prestressing bolts is the T-shaped type of torque spanner turned with both hands (Fig. 17-12). The Tona factory in Czechoslovakia makes such a spanner with a range of 0 to 500 Nm. The tightening force is shown on a dial. The turning moment exerted on the spanner and nut is indicative of the tension generated in the bolt. The moment necessary to start turning the nut at any point is usually 20 per cent, higher than the moment required once the starting resistance (static friction) has been overcome. Universal pneumatic hammer drills (see Fig. 14-2), or pneumatic impact

240

tools (Fig. 17-13) equipped with wrench adaptors can be used to tighten bolts more quickly and easily (see Section 14.1), torques of up to 832 Nm and more being obtainable with these machines.

IMPACT TOOL

Fig. 17-13. Light-weight impact tool (Ingersoll—Rand) with adaptor; torque capacity up to 4,000 Nm

Fig. 17-14. Light hydraulic stressing equipment of Elbroc, Bate man Co.

Torque wrenches and pneumatic impact tools enable a relatively accurately controlled torque to be exerted on the bolt nut, but the tension in the bolt is less predictable because of variability in the resistance to revolution of the nut. Several controlled experiments have shown that the actual tension in the bolt can vary within a range of ±25 per cent, of the stated value [208]. For this reason, there is a tendency to prestress bolts using light hydraulic jacks. In South African mines, for example, a one-man, hand operated, hydraulic jack manufactured by Elbroc, Bateman (Fig. 17-14) is used. This is a compact piece of equipment weighing approximately 10 kg, which incorporates a fixing device that is screwed on to the protruding end of the bolt together with the nut, a hydraulic pump for generating tensions from 10 to 100 kN, and a device for tightening the nut when the bar is fully

241

tensioned. The equipment has no delicate gauges; it automatically indicates the tension setting, and stops at this value. Another type of tensioner of the same make with a built-in pump is shown in Fig. 17-15. a)

Fig. 17-15. Durable rock bolt tensioners of Elbroc (South Africa) a) — Mark VIII (0—100 kN), b) — Mark X (0—300 kN) b)

242 17.6.2

Prestressing of bar anchors of high load-bearing capacity

For the larger bar anchors with higher load-bearing capacities and prestressing forces greater than 100 kN, larger hydraulic jacks have to be used. Various types of jack and hydraulic equipment designed to grip the threaded end of the bar (or cable) are available on the market (e.g. Enerpac). They are generally small in size, but have a high performance (Fig. 17-16). The

Fig. 17-16. Centre hole hydraulic jacks Proceq CP 100 and CP 150 with manual pump for anchor testing and pre-stressing up to the tensile load of 1 MN and 1.5 MN respectively

end of the anchor bar passes through a hollow cylinder in most types. In some cases (e.g. the American CCS system), the stressing equipment consists of a couple of small jacks connected by a bridge to which the end öf the bar or anchor rope is fixed. Weights of the hollow cylinder stressing equipment used by the Dywidag Co. are quoted as an example; the equipment for tensioning bar anchors with forces of 250, 600 and 1,100 kN weighs 23, 36 and 47 kg respectively. Also the equipment used in Great Britain and in other countries is relatively easy to handle. A typical arrangement cf the stressing equipment for bar anchors is shown in Fig. 17-17. Oil is supplied to the hydraulic jack by a pressure pump. A hand-operated pump is sufficient for dealing with a small number of anchors. For larger

243

I

sgroutpad

hydraulic pump {manual)

hydraulic hollow ram-jack

Fig. 17-17. Arrangement of stressing equipment for bar anchors, according to Littlejohn and Bruce [120]

Fig. 17-18. Alevin pump BA 3 (max. pressure 60 MPa, delivery 6.5 1/min)

anchoring schemes, electric pumps (or less often combustion engine pumps) are employed. The stressing equipment is generally designed to generate working pressures of 40 to 70 MPa. Some of the specification details of pumps used by the Stronghold Co. (working pressure, 60 MPa) are given for the sake of illustration. A portable pump with an output of 2.62 litres/min, capacity 20 litres, weighs 90 kg and is 750 mm long and 540 mm in height; pumps with outputs of 3.7 and 6.5 litres/min, capacity 45 litres, weigh 210 kg and 220 kg, respectively. These two pumps are mounted on a simple undercarriage (Fig. 17-18) and difTer only with respect to the driving power. Where the anchor tendon is composed of a bundle of bars, bars are usually stressed one at a time (Fig. 17-19a). In Great Britain a piece of equipment

244

Fig. 17-19. Stressing of Macalloy bar system a) — individual bar stressing, b) — equipment for stressing 4 bars at a time

for prestressing a bundle of four bars together has been developed (Fig. 17-19b); allowance is made for the nuts at the ends of the bars to be easily screwed on and tightened. With this equipment a total prestressing force of 2.2 MN can be applied. It is comparatively light and small, can be easily adapted for prestressing a six-bar bundle with a force of 3.24 MN.

245

When the required tensile force is produced by the jack, the anchor tension is maintained by tightening the nut with a spanner, which gains access to the nut through an opening in the spacing chair below the jack (Fig. 17-20). Perfect seating of the nut on the bearing plate is indicated by a slight drop in pressure on the jack manometer (equivalent, approximately, to 5 kN force at the jack). After this has occurred, the pressure in the equipment can be released altogether.

Fig. 17-20. Stressing of anchors to 500 kN with Proceq Co. equipment. (Note opening in the spacing chair for tightening the anchor nut, using special spanner)

Specially designed jacks for prestressing Dywidag bar anchors are shown in Fig. 17-21. A ratchet wrench for tightening the anchor nut after the tendon has been prestressed is incorporated in the jack. 17.6.3

Prestressing of cable anchors with locking heads

The tensioning of heads provided with a screw thread on the outside is simple; the stressed state is maintained by a ring nut resting on a loaddistributing headplate (see Figs. 16-3,5, 10,21). With this arrangement, no loss of prestressing can occur from slip, and the prestressing may be increased

246 Fig. 17-21. Electrically powered hydraulic jack with built-in socket wrench for single bar anchors (Dywidag Co.). Maximum jacking force, 590 kN

Fig. 17-22. Stressing gun of up to 2 MN formerly used for cable anchors by the Losinger Co. The anchor nut is tightened by hand on to the distribution plate after prestressing

if necessary (see Section 16.1). These anchors are stressed, like the e.g. PZ anchors, by means of a hydraulic jack with a hollow cylinder; the threaded rod passes through this cylinder and is fixed into the centre of the anchoring head at one end and onto the cylinder face by a screwed on nut at the other end. The required prestressing is maintained by tightening the nut or ring on the outside of the head (Fig. 17-22), or by fixing the tendon in position with a nut bearing on the surface of the anchored structure. In the latter case, the rod in fact is an extension of the anchor (see Fig. 16-3). Most anchors with locking heads holding cables made up of individual wires or strands are prestressed with special prestressing guns. The wires or strands are fixed either around the perimeter of the gun on the gun cover (Fig. 17-23), at one end of the hollow cylinder in many cases (Fig. 17-24),

247

./..

nJ% \ m. m,4

I

Fig. 17-23. Anchor head and stressing equipment for 1 MN anchors of smooth patented wire (Horel system)

Fig. 17-24. Hydraulic jack in use for anchor stressing, and anchor head shown after prestressing {Losinger system)

248

inside the hollow cylinder (see Fig. 17-27), or around the perimeter of the jack as well as at the end of the hollow cylinder (Fig. 17-25). On placing the anchor, sufficiently long ends of the wires or ropes must be left projecting. from the borehole for fixing to the stressing equipment. The cable wires or strands are fixed with wedges into a ring strongly fastened around the perimeter of the gun; at the end of the hollow cylinder they are fixed into a stressing head of a design corresponding to that of the anchoring head (Fig. 17-26). The gripping wedges, being re-usable, are made of special steel, and must be lightly greased before they are placed so as to facilitate removal after prestressing has been carried out.

Fig. 17-26. Arrangement of stressing equipment fixed at the anchoring head of a multiple strand cable anchor before it is prestressed (documentation VSL)

Prestressing of cable anchors of strands with locking heads proceeds usually in the following way (see Fig. 17-26): 1. The individual parts of the tendon, anchoring head and bearing plate are thoroughly cleaned. 2. The anchoring head is slipped on to the tendon, and moved along until it touches the distribution plate, by pulling the individual ropes through the openings in the head; the gripping wedges are inserted in the openings, beside the strands. Care must be taken that the strands remain parallel, and do not cross. 3. The resting chair, stressing jack, and stressing head are mounted on

249

the tendon, and lightly greased wedges are inserted into the openings in the stressing head. Prior to the stressing of short anchors, it is important to make sure that the displacement of the anchor head will exceed 30 mm under the highest loading. Otherwise it becomes impossible to remove the gripping wedges from the stressing head following the release of the jack. Wherever a displacement of 30 mm or less is expected, the jack piston must be advanced 30 mm before the stressing and mounting of the stressing head are carried out. 4. The oil hose from the high-pressure power or hand-driven pump is connected to the jack, and the stressing is begun. 5. With the initial movement of the jack piston, the tendon ropes become fixed in the stressing head. The anchor head and its free wedges are held in position by the resting chair. 6. The pressure of the jack is progressively raised to the maximum value required. The magnitude of the pressure and the advance of the piston are carefully observed and recorded. 7. When the prescribed tensile force, or the maximum advance of the piston has been reached, the pressure in the jack is released and the ropes are automatically fixed in the anchor head by the forcing of the wedges into position. The stressing head is freed by light tapping. 8. The jack can be disconnected, or, after retraction of the piston to its original position, stressing can be continued to the next stage by repeating the procedure. To obtain the final tensile force in the tendon, or in compensating for losses of prestress, a higher resting chair and headplates are positioned under the anchoring head. Equipment for prestressing cables varies only with respect to size, thrust distance, and working pressure. Some parameters of BBRV stressing equipment are given in Table 17-111. The Cona-Multi system of the BBR Co. consists of a series of prestressing guns for forces ranging from 110 kN up to 10,000 kN. The equipment for 10 MN (CM 1000), has a cylinder of diameter 725 mm and thrust distance 400 mm, and weighs 3,380 kg. TABLE 17-111 Parameters of BBR V stressing equipment Type

NP60

NP100

NP 200

NP 300

NP 500

force (kN) cylinder diameter (mm) thrust distance (mm) weight (kg)

620 205 100 28

1,030 260 100 85

2,060 290 100 117

3,090 350 100 196

5,150 560 100 1,260

250

The BEB Multiforce Ram 12x12.5 mm system weighs only 15 kg for equipment producing a prestressing force of 320 kN, and 280 kg for 1,920 kN equipment. The West German Eberspächer Co. and Losinger Meili of Switzerland employ also hydraulic jacks with a hollow cylinder. This range of equipment includes types producing forces of 100 to 10,000 kN. The specifications of the 10 MN equipment are: 780 mm diameter, 54.9 MPa working pressure, 1,884 kg weight at a thrust distance of 200 mm, and 2,260 kg at 300 mm. The 10 MN types are designated ZD 869 and ZD 870. The diameter of the internal opening is 260 mm for both types. The English Stronghold system (Fig. 17-27) again uses a hollow hydraulic cylinder but without an opening. It differs from the other types in having the wires or ropes fixed inside the cylinder rather than passing right through the cylinder. With this arrangement, economy in the length of the cable

£^fc^4l

Fig. 17-27. Operating sequence of the Stronghold jack in the stressing of an anchor cable

is achieved. The cable components are fixed in the anchoring head by means of an indexing plate. This template, after release and restressing, presses the gripping wedges into position and seats them forcibly under pressure from the jack's lock-off mechanism. Reversal of the oil flow retracts the jack which automatically releases the internal fixings for removal and preparation for the next operation. The individual steps of

251

the sequence are remotely controlled from the pump without the need for attendance at the jack face. The range of equipment available from this firm includes jacks with forces of 600 to 12,000 kN. The specifications for the 12 MN equipment are: 900 mm diameter, 80 MPa working pressure, weight 2,200 kg, 400 mm thrust distance; this type is designated G-1200. The French Freyssinet system made by the STUP Co. employs a traditional prestressing gun with the wires or strands fixed around the perimeter; in the guns that produce greater forces, the cable components also pass through the gun, and in the latest versions a hydraulic cylinder with a central opening is in use. The weight of the equipment depends on what working pressure is required, the smallest weight being 400 kg and the largest 900 kg {Freyssi Monogroup). Even when high working pressures are applied, equipment for prestressing large anchors is difficult to handle on account of its bulk. Some companies therefore use lightweight equipment designed to stress the individual strands of which large anchors are composed. The Italian Tensacciai system, for example, is based on light prestressing guns for the stressing of individual ropes. It is used for prestressing anchors composed of up to 20 strands each of 15.2 mm diameter, prestressed to a total force for whole anchor of 3.8 MN (working capacity 3.02 MN). The process can be speeded up by using a number of guns together (Fig. 17-28), and for this reason the guns are made with a small diameter and are connected to a central pump. The easily portable Stöbet gun system with an output of 60 to 200 kN is made in Bulgaria under licence from the German firm of Max Paul. These guns weigh only 16 to 20 kg. The Czechoslovakian VUIS system employs equipment designed for twin ropes (Fig. 17-29).

Fig. 17-28. Prestressing of multi-rope anchors by means of a system of light Tensacciai guns

252

βββιβ~

^^Ξ~£^

**«: - \ ^τ^

Fig. 17-29. KC//5 system equipment for prestressing a couple of ropes of 15.5 mm dia

Fig. 17-30. Head with immovable fixing of tendon a) — cross-section of a head showing cushions (supports) made of steel pipe filled with concrete b) — view of a prestressed head with supporting steel plates

253

17.6A Prestressing of anchors with immovable head fixings Where these anchoring heads are provided with an external thread (e.g. the SEEE and BBRV systems), they are stressed in the same way as bar anchors or threaded locking heads. Tension is maintained by tightening a nut. For the more robust anchor heads in which ropes or tendon wires are fixed with grout (Fig. 17-30), individual stressing equipment was used. It was equipped with currently used hydraulic jacks (Fig. 17-31). The prestressing is maintained by means of headplates under the anchor head (see Fig. 17-30). The stressing of an anchor of this type with a force of 10 MN on the Bou-Hanifia Dam in Algeria in 1936 is shown in Fig. 17-32. Various designs of these heads are described in the Section dealing with the anchoring of gravity dams.

Fig. 17-31. Stressing of an anchor with a force of 4 MN VUIS. Dynamometers are placed under the stressing equipment for checking the induced force

254 Fig. 17-32. Prestressing of anchoring cables by a force of 1,000 metric tons (10 MN) on the Bou-Hanifia Dam in Algeria in 1936

h*tf* f]

v1, 4* / **-v;

i J ' ^H

ri| «*>· ' ,; **W * i " ..Γ* v ^ - /v l' '"if; *" >, "*^*χ% . ?Wv~

■ * * * > .

-

* * i \

^ ;'i *

s

V ■

•Sf

^

-<

^

'*<€'

Chapter 18 P R O T E C T I O N OF A N C H O R S A G A I N S T

CORROSION

The service life of structures anchored in rock depends upon the durability of the anchorage. The greatest threat is from corrosion, which is particularly likely to occur in the environment in which the anchor root is embedded.

18.1 PRINCIPLE OF CORROSION

Corrosion is the damage caused to metals by their chemical or electrochemical reaction with the surrounding medium, typically heterogenous reactions taking place at the boundaries between solid, liquid and gaseous phases. The mechanisms of corrosion are governed by many factors which cannot be easily defined, and which, moreover, change in the course of the reactions themselves. It is thus extremely difficult to give a reliable explanation of corrosion; simplified accounts are often presented, but these cannot be considered as being comprehensive or reliable. It is safer to draw conclusions from the results of experiments wherever anticorrosive treatment is planned, particularly in large projects. It may be stated generally, that steel structures embedded in the ground, mainly suffer the type of corrosion caused by electrochemical reaction; this entails a transformation of the metal into free ions and electrons as a result of the interaction of the metal surface with an electrolyte (soil moisture). If ions of only one metal take part in this reaction, the process is anodic, the ions passing from the metal into solution as free hydrated ions; when metal ions separate out from the solution and recombine with the metal, the process is cathodic. Corrosive processes involve both anodic and cathodic reactions. They are usually limited to definite areas on the metal surface owing to the heterogenous nature of the metal, or because of differences in the composition of the ground in which the metal is placed. It is the electric current passing between these specific areas which causes the corrosion. The cathodic reaction reinforces the anodic reaction by drawing off the electrons released by the latter. After some time, however, the primary anodic and cathodic products mutually combine to produce insoluble substances which prevent further corrosion, unless oxygen or hydrogen arising from the reduction of hydrogen ions penetrates the metal at the site of corrosion. These products disturb the

256

anodic and cathodic processes and facilitate further transformation of the metal into the free ion form. The ground consists of solid, fluid, and gaseous matter, and nearly always provides suitable conditions for electrochemical corrosion. Even the slightest moisture within the ground capillaries can act as an electrolyte, and it is not necessary for potential anodic and cathodic sites to be completely surrounded by electrolyte. In fact, metal corrosion tends to be depressed in saturated rocks (Fig. 18-1). The degree of aeration, and its local differences have a considerable influence on the progress of corrosion. Thus, for example, the anodic part of the corrosive process may take place in an area surrounded by unaerated medium, yet is accelerated if other areas are exposed to aerated medium. 1.2

Ί.ο

A

>/ / /

0.6

0Λ 0.2 / 0

>

y

L·

0

4

Λ

)

tΎ

z> \ 9 _A

3

*=q

8 12 16 20 24 humidity [%]

Fig. 18-1. Effect of soil moisture content on the corrosion of steel 1 — clayey-loamy soil, 2 — sandy-loamy soil I, 3 — sandy-loamy soil II

Under some conditions the metal of anchors is also attacked by aerobic and anaerobic bacteria. This biocorrosion can be caused, for example, by an irregular distribution of bacteria on the metal surface, which produces varying degrees of aeration. Bacteria also accelerate chemical processes, such as the formation of sulphuric acid from pyrites. The chemical composition of the soil has an important bearing on the development of corrosion. The presence of salts, particularly those of sodium, calcium and magnesium (generally acidic carbonates, sulphates or chlorides), particularly encourages corrosion. Because of their high solubility, these salts are mobilized readily by the soil moisture. In order to select the best method for protecting anchors against corrosion, the corrosive activity of the ground in which the anchors will be embedded must be known. The corrosive activity is determined by making measurements of the following: a) the composition of the ground and the ground water level; b) the virtual ground resistivity; c) the specific conductivity of the ground water and surface water; d) the chemical composition and moisture content of the individual geological beds of the ground;

257

e) the chemical composition of the ground and surface water; f) physical and chemical properties (pH, redox potential, etc.); g) the possible existence of extraneous electric fields. When the anchors have been fixed, it is expedient to check the anchor/ ground potential by measurement. A summary evaluation of the aggressiveness of the rock, soil, and water to steel, according to the results of the investigation, can be worked out from Standards which are often available for the evaluation of the aggressiveness of ground conditions to steel pipes. It is essential that at least two of the data from Table 18-1 should correspond with measured values to state the respective degree of classification. If the anchor tendon passes through beds of different geological type and composition, the danger of corrosion is greatly increased. TABLE 18-1 Classes of aggressiveness of ground to steel Aggressivity of the ground

Virtual ground resistivity [Ω/m]

Water conductivity [^S/cm]

Redox potential [mV]

I low

> 100

< 100

400

II medium

50 to 100

200 to 100

200 to 400

III high

23 to 50

430 to 200

100 to 200

IV very high

< 23

> 430

100

TABLE 18-1 cont. Current density in ground [mA/m2]

pH

Rock or soil content of total sulphur [%1

< 1.10"4 3

MO"

to3.10"

> MO"

1

Cl [%1

so3 + ci

co 2

[mg/1]

[mg/1]

6.5 to 8.5

< 0.1

< 0.02

< 100

0

4

8.5 to 14

0.1 to 0.2

0.02 to 0.05

100 to 200

0

3

6.0 to 6.5

0.2 to 0.3

0.05 to 0.1

200 to 300

5

< 6.0

> 0.3

0.1

> 300

5

3.10" to l.lO" 1

Water content of aggressive

258 The cables from which anchors are made, consist exclusively of wires treated by patenting and cold drawing. The high mechanical stresses in the individual strands of a rope can contribute to the development of severe and very dangerous corrosion under stress, including corrosive cracking and hydrogen brittleness. These corrosive effects develop into intercrystalline and transcrystalline corrosive attack. It is in the nature of these types of corrosion that they are hardly noticeable at the outset; subsequent stages, characterized by very fine cracks without any visible products of corrosion (rust), can be discovered only with the aid of a microscope. Failure of the wire occurs at once, without any preceding drop in strength. In this context, it should be noted that of all the processes employed for heat-treating wires, patenting is the most suitable.

Fig. 18-2. Cracks caused by corrosion in a steel wire viewed under the microscope (cracks invisible to the naked eye)

Patented wires are less susceptible to corrosive cracking than wires heattreated by other methods. The danger of corrosion under stress is much reduced by a further treatment following patenting and drawing, namely low-temperature tempering and curing of the wire. This process removes a substantial part of the non-uniform internal stress, which in turn reduces the total mechanical stresses developing within the wire as a result of the summed internal stresses (strain), and those stresses induced by external forces (prestress). Great care must therefore be taken to protect anchors

259

against corrosion, and every design for a permanent anchor in rock or soil must include details of efficient protection measures. Anticorrosion measures for anchors are considered, with respect to three main criteria: 1. Method of providing anticorrosive protection: a) Direct protection by coating, wrapping, or sheathing with waterproof material which keeps out the aggressive external medium. This method is also referred to as passive protection. b) Electric cathodic protection by creating an electric circuit, the anchor surface thus being cathodically polarized and maintained at a potential which prevents occurrence of the corrosive process. This method is designated active protection. It should be noted that cathodic protection is usually applied in the form of a complementary installation for the protection of a large group of anchors; the method is discussed in later Sections. 2. Anchor type: a) Permanent protection such that the service life of the anchor corresponds to the service life of the anchored structure. b) Short-term protection of subsidiary site anchors in service for not longer than five years. 3. Requirements of the static function of the anchor: a) Anticorrosive treatment which allows for static co-operation of the anchor with the ground along its entire length or a part of its length, which nearly always includes the full root length. b) Anticorrosive treatment which prevents adhesion of the anchor to the ground within a predetermined section of the anchor; usually the tendon receives this type of protection. Obviously the designer of the anticorrosive treatment must take into account the functions of the individual anchor parts that are to be protected, and specify the details of the treatment accordingly. For the majority of anchors, different anticorrosive treatments for the tendon and root sections must be considered. Anticorrosive protection must also be provided for the third main part of the anchor — the head. The treatment here is comparatively simple, because the heads are easily accessible and their condition can be checked directly. A special type of treatment is involved in the temporary protection of anchors; some degree of protection must be given to newly made anchors which are awaiting installation, or to the material (wires, ropes, steel bars) from which anchors are to be made. This protection serves during the transport, storage, and handling of anchors, until its function is eventually taken over by the more permanent treatment specified in the design.

260 18.2 DIRECT PROTECTION OF ANCHORS AGAINST CORROSION

Direct protection means that the steel is safeguarded without attempting to remove the causes of corrosion. This protection is designed according to the purpose for which the anchor is intended, the type of corrosion expected, and the insulating material available for making a protective enclosure. The design of any sort of structural protection should be accompanied by recommendations for carrying out the work efficiently, taking into account that the protective material will be applied to the anchor under difficult conditions, without any possibility of exact control of the process. The latter problem has lately given rise to the practice of insulating anchors in special workshops, where the anchors can be assembled and conditions created for closely supervised preparation of the protective sheeting and other materials. Further protection carried out at the installation site is then additional to the primary treatment. 18.2.1

Anticorrosive protection adhering to the anchor

Anticorrosive protection which adheres to the anchor and provides for static co-operation with the ground, is used mainly for short anchors (bolts) of small loading capacity intended for the stabilization of the rock faces of underground caverns, slopes, etc. It is generally used without reference to the rock type for the root section of the anchor, on account of the static function of the latter. The material used to envelop the anchor both as an anticorrosive protection and as a fixing in the ground, is in most cases grout, although synthetic resins are occasionally used (see Chapter 13). From the point of view of anticorrosive protection, cement mortar (grout) is preferable. According to its composition, however, hardened grout is itself subjected to direct corrosive effects of the medium. It is threatened particularly by rock with a high content of sulphate. The rate of corrosion is much affected by the moisture content, the degree of aeration of the ground, and by local gradients of aeration. Since the grout itself cannot be protected from access by aggressive agents, the harmful effects have to be resisted by adjusting the composition of the grout. The activation of cement in activating mixers, the use of plasticizing additives which reduce the water/cement ratio, stabilizing mixtures, etc., all help to build up the resistance of the grout. Hardening accelerators, however, must be avoided, because all types in use at present increase corrosion, quite considerably in most cases. When these accelerators come into contact with a moist ground, they give rise to very strong electrolytes which induce electrochemical corrosion. The grouting pressure applied is also very important, because the degree of compactness and impermeability

261

of the enveloping concrete depend on it. By increasing the pressure the pores and joints in the rock surrounding the borehole are filled with greater certainty, the rock is better sealed, and access of water to the anchor wrapping proper is prevented. A compact cement wrapping 3 to 4 cm thick around an anchor gives fully reliable anticorrosive protection. The long-term anticorrosive effect is determined by the alkalinity of the cement wrapping. Furthermore agents which create an inert layer of calcium ferrite and other calcium salts on the surface of the anchor bars or wires, are released from the cement in the course of tricalcium silicate hydration. For this reason it is preferable to use Portland cement for grouting anchors. Mixed cements are less suitable, because they create a lower degree of alkalinity in the medium surrounding the anchor. It appears from the diagram of pH and electrical potential for iron, that this metal is best protected from corrosion at pH values between 9.8 and 12.3. Hence under normal temperatures, corrosion of an anchor enveloped in an integral body of cement mortar cannot occur, even if the capillaries of the concrete in contact with steel contain water; this is because aqueous solutions of the hydrated products of cement have a basic reaction between p H l l and pH 12.5. When the anchor is stressed, the steel components elongate. It has been demonstrated that the stress in the free anchor section is transmitted into the grouted section for a distance equivalent to 0.3—0.6 of its length; hence, cracks tend to appear in this particular region of the grouted section, that is, in the root part of the anchor. The width of the cracks depends on their density along the surface of the steel, while the density of cracks depends on the grout/steel bond and the characteristics of the ground in which the anchor is fixed, particularly the jointing of the rock. An increase in the density of cracks and the resulting narrowing of these cracks to an admissible value, can be achieved by using bars with transverse ribs or a pressed-on thread (as employed by the Dywidag Co.), or by dividing the anchor root into the largest possible number of separate wires in order to increase the surface area of the root in contact with the grout. The adverse effect of the presence of joints in the ground can be partly eliminated by effective grouting, and if large cracks are liable to appear, these can be filled by repeated grouting after the anchors have been prestressed. Collared pipes are used to enable the grouting to be repeated (see Chapter 13). 18.2.1.1 Double protection of anchors with reliable ground fixing Where anchorage is exploited as a long-term permanent stabilizing element, additional protection with plastic sheets of synthetic dielectric, and waterproof material (polyethylene, PVC, etc.) has increasingly been applied in

262 recent years. This additional protective layer has an indented profile which is acted upon by the shearing forces between the steel and the grout and ground (Fig. 18-3).

Fig. 18-3. Geotest anchor system protected in the root section by plastic corrugated (concertina-like) ducting

The length of the (concertina like) corrugated pipe is determined by the length of the anchor section which is intended to co-operate with the rock, that is, the root length in most cases. Usually the anchor ropes or bars are protected individually in pipes of small diameter; less often, as for example in the BBRV, IRP-Tirsol or APS Tensacciai systems, the entire bundle of ropes forming the anchor is protected by a single common pipe (Fig. 18-4). In the first case, the anchor ropes or bars are grouted into the protecting pipe in advance in the workshop where the anchors are prepared. The grouting can be carried out with special equipment, and the filling of the pipe checked. Wires or ropes grouted in the pipes remain flexible enough to be undulated by means of alternating spacers and clamps, thus giving a greater fixing strength of the anchor in the borehole (Fig. 18-3). Laboratory tests together with experience gained at anchor sites indicate that a pipe diameter 2 to 4 mm greater (at its narrowest point) than the diameter of the enclosed steel component, suffices for reliable envelopment of the steel with grout, and gives a strong fixing of the anchor. The free space between the steel and the pipe wall is then only 1 to 2 mm wide. The grout must be prepared with very finely powdered cement, and a plasticizing additive to increase the flexibility of the mix. Where a corrugated pipe of large diameter protects a whole bundle of strands or bars, the inside space is grouted after the anchor has been inserted into the borehole (Fig. 18-4).

263

b)

o)

Det

^N

4 "ΓΚΤ

KSO^IIIJ

,■

r

/ / ML^VM /

III L

J-"-*"1

11

I

IjxsS

Fig. 18-4. Root of a multiple rope anchor protected by a common corrugated pipe a)—with the possibility of regrouting inside the sheath, b) — with the possibility of regrouting both inside and outside the sheath (APS Tensacciai system) 1 — grip, 2 — anchor plate, 3 — smooth PVC sheathing, 4 — manchette valve, 5 — resin pad, 6 — obturator bag, 7 — corrugated PVC sheathing, 8 — strand, 9 — end cap, 10 — distance piece, / / — injection tubes

264

Collared pipes are used for grouting so that this can be repeated after the first grouting of the space between the pipe and the borehole wall. The space between the anchor sheath and the borehole wall is usually filled with grout, although synthetic resin is occasionally used. Interlocking between the anchor and the concrete is ensured on account of the corrugation of the pipe. The grouting pipes are inserted as far as the end of the anchor root. A vent pipe is also lowered as far as the upper end of the section that is to be grouted, so that air and water can escape from the borehole. Both pipes pass through an obturator (bag) which makes it possible to grout the section under pressure. When the borehole is full, the vent pipe is closed to allow pressure to build up; this pressure is essential for compaction of the grout and penetration of the grout into the spaces in the ground around the borehole (see Section 13.2). The pipes used for grouting are made of rubber, polyethylene or other plastic material, because these are easier to handle than steel pipes. The grouting pipes are inserted together with the anchor; where cables are used, the pipe is sometimes passed down the centre of the cable. In some systems, grout pipes are not used at all, the grout being pumped along the entire borehole cross-section, or into the casing (see Section 13.2). In short vertical boreholes, the grout may also be poured in, provided the holes do not reach below ground water level. Secondary grouting, for the purpose of filling small cracks in the concrete wrapping of the root, is accomplished with the aid of a collared tube. In some instances, cracks in the concrete filling of the borehole may be sealed by pumping grout into the root zone through extra boreholes. 18.2.2

Anticorrosive protection of anchor tendons

In nearly all anchors, with the exception of bolts fixed in sound rock, perfect freedom of movement of the tendon (the section between the head and the root) in the coaxial direction must be established, otherwise the anchor cannot be prestressed. Sometimes the prestressing of fixed anchors decreases due to creep of the ground, and must be restored to the original value; or the ground may be additionally loaded by the weight of a newly built construction, and the anchor prestressing must therefore be reduced proportionally (see Chapter 7). For such adjustments to be made, the tendon must be able to move freely. Displacements of the anchor tendon are made possible by the spreading of insulating layers on the anchor surface, or by locating the tendon inside plastic or metal pipes (see Figs. 18-4, 18-11, 18-13).

265

18.2.2.1

Free insulating layers for anchor tendons

Insulating layers must be correctly applied and must have the following properties: they should prevent the access of moisture to the anchorage; they must be pliable and abrasion-resistant so as not to become damaged during manipulation of the anchor; they should be non-conductive, thus preventing the formation of galvanic cells on the surface of the steel, or they should, by their nature, give rise to a potential at which corrosion cannot ensue; they must be durable and resistant to chemical attack; finally, they should be sufficiently thick and plastic to allow displacement of the internal surface adhering to the anchor tendon relative to the external surface adhering to the borehole wall (or the concrete in the space between the borehole wall and the tendon). It is important in any case, that no displacement should occur between the anchor tendon and the inner surface of the insulation with which it is in contact. The insulating coats are made from various bitumenous materials reinforced with a protective fabric, or interlain with plastic membranes. The thickness of the bitumenous protecting layer is usually 3 to 10 mm, but in some cases may be more. The surface of the protecting layer is generally shielded from mechanical damage by a 10—15 cm-wide insulating bandage made from Polyvinylchloride, polyethylene, or impregnated glass fibre fabric. The bandage is wound on to the anchor with a half width overlapping, and is glued on to form a single integrated wrapping. Until recently, insulating coats were made with asphalt modified by an admixture of plastics. These materials afforded a very high degree of insulation provided that they were enclosed to prevent viscous flow of the material and that the asphalt was washed with running water. To prevent flow of the insulating layer, it must be as viscous as possible, and protected by a covering that can withstand mechanical damage. The asphalt coat of the anchoring cables used in the reconstruction of the Cheurfas Dam (1934), was protected by an impermeable cloth wrapping. The insulation was applied as the cables were lowered into position, by pouring molten asphalt into the wrapping and progressively closing the wrapping with a zip fastener. Uniform thickness of the asphalt layer was obtained by winding a rope in a helix around the cable, underneath the wrapping (Fig. 18-5). In recent years, the development of insulating materials has seen extraordinarily rapid progress. The new insulating materials which form the coat not only form a barrier preventing access of the corrosive medium to the urface of the anchor, but also contribute towards changing the properties of the corrosive products and increasing the protection of the anchor by being electrochemically active, and by being able to inhibit corrosion by

266

changing the composition of the surrounding medium. Agents are added which considerably reduce the effects of those aerobic and anaerobic bacteria and moulds causing biocorrosion. The new insulating materials, when correctly selected and applied, guarantee a long-term, almost unlimited, service life for the anchor.

Fig. 18-5. The insulating wrapping oF cables used in thefirstreconstruction of the Cheurfas Dam 1 — hemp rope, 2 — asphalt filling, 3 — watertight wrapping closed by a zip fastener, 4 — borehole wall

The best procedure is to apply an insulating coat of anticorrosive paste combined with insulating bandages. The primary insulating coat of anticorrosive paste is particularly recommended for anchors consisting of ropes. The paste fills the spaces among the strands nearest the surface of the rope. These spaces do not alter shape when the ropes are flexed, transported, or inserted into boreholes; this is not the case, however, with cables assembled from individual wires. The paste spread on the anchor surface and wrapped in the appropriate bandage, forms a condensed unit which is plastic enough to allow for movement of the anchor during prestressing; at the same time it is compact enough to prevent access by corrosion-inducing agents to the surfaces of the anchor wires. In Czechoslovakia, a combination of PLU anticorrosive paste with bandaging is employed. The PL U paste consists of a basic paraffin substance containing 50 to 60 per cent, mineral matter, particularly barytes and glass, and 5 per cent, alkaline chromate sludge. A very important characteristic of the anticorrosive paste is that in the presence of water the chromates are extracted from it very quickly. Hence, if water penetrates the insulation (the latter, for example, having suffered mechanical damage), or if the paste has been spread on to wet steel compo-

267

nents, the inhibitors are rapidly extracted at high concentration. The anticorrosive PLU paste is also a strong fungicide because of the presence of chromates. The PLU bandage is made of glass fibre fabric, 50, 100 and 200 mm wide, treated on both sides with a compound of similar composition to that of the paste. The biological resistance of the insulating layer is enhanced by the presence of heavy metal salts and cyclohexane-carboxylic acids. The amount of compound used is 1.60 kg + 5 per cent, per 1 m 2 of the fabric. The PLU paste and bandage were used for the first time in 1957, and insulation made at that time with these materials has retained its original properties to this day. Hence, assertions as to its long service life are fully justified. Another similar product is Plastikor. The fabric of this insulating bandage is made as Arachne non-woven propylene textile; it is impregnated with a special plastic anticorrosive compound containing corrosion inhibitors, polar substances and fungicides. Its main advantage over the PLU bandage is its greater elasticity, which makes it easier to use, and more effective on uneven surfaces. The general procedure for applying insulating bandages is relatively simple. The PLU or Plastikor bandage is tightly hand-wound in the cold state, with 50 per cent, overlap, on a clean anchor cable, or better still, on a cable coated with anticorrosive PLU paste (Fig. 18-6). The bandage is smoothed by hand or with a special tool so that it forms an uniform layer. A well made wrapping should be at least 3 mm thick. The thickness may be checked non-destructively by means of eddy currents —the principle used, for example, in the portable Isotron Fe apparatus. Another convenient test is the so-called spark test, which helps to find any deficiencies in the insulation such as excessive porosity, insufficient thickness, and mechanical injuries invisible to the naked eye. The insulation should resist puncture by a spark of up to a minimum of 15 kV. To avoid mechanical injury while inserting an anchor into the borehole, the insulating bandages can be protected by wound-on PVC strips. The cables of the Firth of Forth suspension bridge in Scotland are insulated by this method. They are composed of 60 patented 5 mm-diameter wires insulated by a triple coating. The first layer is formed of Denso insulating paste which was spread on the individual wires on a work-table. The basic component of the Denso paste is a hydrocarbon derivative of paraffin oil, with a siliceous filler and anticorrosion agents. This paste removes all surface moisture in contact with the steel, and neutralizes oxides which have formed. It does not harden, and permanently retains its insulating properties (the insulation is guaranteed for 100 years). The second layer consists of a Denso bandage, wound with 50 per cent, overlap and smoothed to form an impermeable wrapping. This bandage is a cotton fabric impregnated from either side

268

"'"**$/: *~~ v«^v

^Jf|WMt,

<φ

^^ * s^S&^fe?· .- . » 4 .

^^ί

1

MB^ISefc: :.Sjf%

Fig. 18-6. Application of insulating wrapping to a cable

269

with a mixture similar to the Denso paste, but with stronger anticorrosive properties. The outer layer provides protection from mechanical damage and consists of a Denselt bandage wrapped around the cable. This bandage is of jute fabric impregnated with plastic^zed asphalt containing an inert filler. The bandage is warmed with a blow lamp as it is wound, to soften the asphalt; on cooling, the bandage becomes a strong impermeable wrapping resistant to abrasion and other possible mechanical damage. The practice of applying insulating wrappings to anchors composed of parallel wires has the disadvantage that it is necessarily carried out at the installation site. Otherwise, the technique is highly advantageous for this type of anchor, since the insulating compound fills the spaces between the wires inside the cable. This is not so with rope cables, in which the interior surfaces of the wires are often attacked by corrosion. Nevertheless the use of rope cables is profitable, because they reduce the labour costs at the site, have a higher ductility which reduces losses of prestressing on permanent deformation of the prestressed rock (see Chapter 11), and, furthermore, they can be insulated in the workshop. To prolong the life of rope cables it has been proposed that the insulating envelope be made in two stages. In the rope factory, the individual strands of the rope are insulated prior to stranding by passing them through a bath of fluid insulating paste. In this way all the interwire spaces are reliably filled. In Czechoslovakia, the rope factory at Bohumin supplies ropes in which the internal spaces are filled with a red lead sealing compound. When the prestressing of such ropes was checked after 16 years in service, no weakening due to corrosion was found, although a short section of the ropes was in contact with humid air; the outer asphalt wrapping was not protected by concrete, but only by a glass fibre bandage in this section (Fig. 18-7). The external insulating layer may also be made in the factory by pulling the rope through a nozzle, the insulating compound being forced into the nozzle at the same time, either as a hot melt or a cold solution. It should be pointed out that the external layer usually cannot replace the function of the primary coat, (of anticorrosive paste, grease etc.) because the applied compound cannot penetrate the spaces in strands of more than 7 wires when they are arranged in several layers. Unfilled spaces admit the entry of water and oxygen, and thus anchors can be attacked during storage and during installation before the insulating wrapping of the tendon has been watertightly sealed at the anchor head and root. Most anchoring systems, however, use seven-wire strand as the basic unit of assembly, and the interwire spaces become filled with the compound without any difficulty (Fig. 18-8), provided that the paste is pressed on to the surface of the strand and thus forced between the wires. In the construction of the London Thames Flood Protection Scheme (Fig. 18-9) the anchor cables used to stabilize the

270

o) b) Fig. 18-7. Anchoring rope 37 x 19 dia 2.9 mm, denuded after 16 years of service a)—superficial damage to the sheath occurred at the level of the ground surface,/?)—surface of the cable stripped of insulation <4 Fig. 18-8. 7-strand ropes and insulating compound filling the internal space of the protecting pipe

Fig. 18-9. Double protection system with encapsulation of the tendons submitted by VAC. (International Construction 1979/Sept.) / — polypropylene-covered LR Dyform strands, 2 — polyester grout, 3 — corrugated lead-coated steel duct, 4 — seal, 5 — recess grouted after stressing of anchor, 6 — anchor cap, 7 — grease

271

quay wall were protected from corrosion with 2 mm-thick polypropylene sheets supplied by Chemical Products Ltd. Applying insulating coats by spraying or brushing is not reliable enough. 18.2.2.2

Protection of anchor tendons by pipes

The application of insulating coats, whether this is done on site or in the workshop, is laborious, and the quality of the work depends on working conditions which cannot always be guaranteed. For this reason, increasing use is being made of a method in which the anchor tendon is inserted into a rigid or flexible plastic pipe. Smooth plastic ducts, made of so-called branched polyethylene, are most often used for this purpose. They are flexible and are supplied in coils in lengths of 50 to 100 m, diameter 60 to 70 mm. Pipes of larger diameter are mostly made of so-called linear polyethylene. The latter pipes are rigid and are supplied in lengths of 6 to 10 m. The spaces around anchor tendons inserted into these smooth ducts are filled in special workshops with a mixture of asphalt and plastics or with an other compound of guaranteed plasticity (see Fig. 18-8), so that the anchor is free to move during the prestressing. When the required additional stressing of the anchor has been completed shortly after its insertion into the borehole, the anchor enclosed in its duct can be grouted on the site. The anchors securing the underground structures of the tyre factory at Otrokovice (Czechoslovakia) against displacement by upward hydrostatic pressure, were protected in this way. For the insulating pipe a flexible vacuum hose reinforced with metal rings was used; this hose enclosed a free space around the tendon, while the space between the hose and the borehole was grouted in order to fix the anchor root. After the prestressing and subsequent additional stressing of the anchor had been carried out, the space inside the hose was grouted by means of a grouting pipe reaching to the lowest point of the hose (in this case close to the anchor root, Fig. 18-10). The grouting pipe was connected to the hose and was inserted together with the hose into the borehole (Fig. 18-11). In the upper section of the space inside the hose, that is, under the anchoring head, a bleed pipe led out of the hose. An anchor should not be left longer than a month inside an unfilled protecting pipe. This period can be extended, however, if a moisture adsorbing agent is placed in the pipe, or the surface of the anchor is provided with a temporary protective sheath (see Section 18.4). Methods have been proposed of providing the anchor wHh a sheath affording its permanent protection against corrosion before it is inserted into the protecting hose or pipe and enveloped in grout (Fig. 18-12). Such an arrangement represents treble or quadruple protection against corrosion; since it is costly and complicated to set up, it is recommended

272 section ΙΊ'

Fig. 18-10. Anchor protected by an insulating tube connected to the steel surround of the base 1 — anchor cable, 2 — insulating tub?, 3 — grout pipe, 4 — borehole, 5 — anchoring cavity, 6 — anchor base constructed prior to installation of the anchor

9;<\$fW

Fig. 18-11. View of an anchor protected by a flexible tube to which a grout pipe is connected for secondary grouting of the tendon

only for the anchoring of very exposed structures. It was used in the construction of the pumping station for the water supply of Bratislava (Czechoslovakia) (see Chapter 22). The wires of the anchor cables were coated with red-lead in the factory. The second insulating envelope consists of a layer of PL U insulating paste and anticorrosive bandage, protected from mechanical injury during manipulation of the anchor by a bandage of PVC foil. The anticorrosive protection is further increased by the casing (diameter 133 mm, wall thickness 12 mm) which remains in the borehole; the space between the casing and the anchor is grouted. Uniform thickness of this grout layer was achieved by winding a 10 mm diameter wire in a helix around the anchor,

273

Fig. 18-12. Design of the permanent anchors used to secure the pumping station supplying water to Bratislava 1 — 2 MN VUIS anchor head, 2 — protective wrapping, 3 — flange for fixing rubber sleeve, 4 — de-aerating pipe, 5 — washer, 6 — steel plate (720 x800 mm), 7 — rubber sleeve, 5 — insulating wrapping, 9 — pipe (135 mm dia.), 10 — casing (135 mm dia.), 11 — cable (65 mm dia), 12 — sealing ring, 13 — load-distributing sill, 14 — sealing for the casing, 15 — walls of the pumping station, 16 — spacing insert of a 10 mm dia. helix, 17 — PL U paste, IS—PLU bandage, 19 — PVC bandage, 20 — cement grout

thus allowing free passage for the grout. At either end, the casing is freely inserted into a pipe, the upper pipe being embedded in the concrete of the seating sill positioned on the surface of the station or caisson wall (see Fig. 18-12). A design has also been used in which the space between the protecting pipe and the cable (the latter bsing provided with an insulating wrapping) was left without an insulating filling. In this case the protecting pipe had to be corrosion-resistant, that part of the anchor tendon inside the pipe had to be provided with a reliable anticorrosive wrapping, and the ends of the protecting pipe, where the tendon was connected with the root and with the anchoring head, had to be sealed thoroughly. The pipe in this case merely played the part of a dielectric. Such an arrangement may assist in the removal of the anchor, if this should be deemed necessary at some future time. 18.2.3

Waterproof connection of the insulating envelope with the anchor root

Every anticorrosive system has some weak point, namely the connection of the insulating wrapping of the anchor with the anchor root being a potential point of failure. This connection is relatively reliable in those cases in

274

which the root is protected from corrosion by a transversely corrugated pipe, because such a pipe can be watertightly connected with the insulating envelope of the tendon (see Fig. 18-4). An equally reliable type of connection is that formed between the insulating envelope and an anchor base with a cylindrical steel sheathing; the insulating wrapping is slipped over a flange on the base and the connection is made watertight with a locking sleeve or bandage (see Fig. 18-10, 18-11). Where anchors are grouted along a section delimited by a seal or bag, reliable grouting of the connection point of the covering with the anchor root and proper filling of cracks which have formed in the root concrete during the prestressing of the anchor (see Chapter 13), is achieved by repeated applications of grout using collared pipes (see Fig. 13-40). The borehole space around the root and the space around the covered anchor are not to be filled with grout in one operation, because the connection between the covering and the concrete of the root cannot be resealed, should it break. The filling of the borehole in two stages, first in the fixing of root zone, and then in the insulated, tensioned part of the anchor, is undoubtedly safer, particularly if a facility is created for re-grouting. Two-stage grouting and re-grouting both raise the costs of anchoring operations. However, these procedures may not always be necessary, provided that the tendon is able to move freely in the insulating envelope during prestressing and that no tension is induced in the grout filling above the anchor root, but rather a compression occurs which adds to the watertightness of the insulation. When the cable is pulled against the restraining influence of the root during prestressing of the anchor, it slips inside the insulating wrapping of the tendon section and remains protected against corrosion. In spite of this, it is good practice to strengthen the insulating wrapping at the transition zone between the fixed and tensioned parts (Fig. 18-13). With a view to improving the protection of the anchor at the neck of the anchoring bulb, it is expeditious to increase the thickness of the insulating layer to the full width of the borehole. When the anchorage is tensioned under these conditions, the movement of the bulb, and the reaction of the concrete filling of the borehole, cause the insulating material to be pressed into any cracks which appear in the concrete of the transition zone during the first stressing stage. This effect can only be realized in anchors with bases. A non-compressible plastic material is applied in a layer 1 to 3 cm thick on the front of the base, and is then covered with a freely mounted cover (Fig. 18-14). The protection of the neck of the anchor bulb is comparatively simple; a metal sheet cover is mounted on the anchor. The effect of this protecting cover is increased by the application of a plastic insulating paste around the anchorage, where it is attached to the bulb (i.e. around the neck). A fully

275

Fig. 18-13. Strengthening of the insulating coat at the transition zone between the tendon and the base a) — in a bar anchor, b) — in a cable anchor; 1 — anchor base, 2 — insulating wrapping, 3 — strengthening of insulating wrapping, 4 — concrete filling of the borehole, 5 — nut, 6 — mounted insulating tube

Fig. 18-14. Diagram showing mechanically compressed protection of the connection point between anchor tendon and base 1 — anchor, 2 — anchor base, 3 — nut, 4 — incompressible insulating plastic material (e.g. PLU paste), 5 — cover, 6 — insulating sheath. 7 — cement mortar

watertight covering of the anchor base is ensured by incorporating the end margin of the insulating wrapping into the material of the anchor base. Such an arrangement was used, for example, in the above-mentioned construction of the Thames quay wall. The tendons, enclosed in a steel pipe, consisted of three prestressed strands insulated with polypropylene sheets 2 mm thick. The ends of the strands were grouted inside the steel duct with a pre-mixed polyester compound. They were inserted individually in such a way that a margin of the polypropylene cover 100 mm wide was also immersed in the grout (see Fig. 18-9). The embedding of a section of the insulating wrapping of an anchor within the base undoubtedly guarantees a watertight connection between base and tendon. However, it reduces the effective distance along which a bond is formed between the anchor wires or ropes and the base filling (i.e. the stress transfer region is reduced); this means that the length of the anchor base must be extended by the width of the margin of embedded insulating cover. From this point of view, it is less costly to protect anchors by the arrangements described earlier, viz. by applying a layer of insulating material (paste) where the cable is inserted into the front of the anchor base

276

(see Fig. 18-14), or by making a watertight connection between the insulating cover and a flange on the base (see Fig. 18-11), or between the insulating cover and an anchor root sheath made of crimped plastic pipe (see Fig. 18-4). 18.2.4

Anticorrosive protection of anchor heads

The protection of anchor cables around the point of connection with the anchor head is easier to achieve, because this region is usually accessible, allowing the arrangement to be checked. Even so, defects in this part of the cable have been known to occur leading directly to anchor failure, and therefore a discussion of design principles relating to the protection of anchor heads, together with some examples of arrangements which have been used, will be appropriate here. Until now, anchoring heads have usually been protected by grouting the mouth of the borehole following prestressing of the anchor, concreting the space under the head base (Fig. 18-15), and eventually, covering the entire

Fig. 18-15. Supporting plates embedded in concrete under the base of the fixed head of a 4 MN anchor. Grouting pipe visible in foreground (Hricov Dam, Czechoslovakia)

head with concrete (Fig. 18-16). These methods, however, do not guarantee effective covering of the bare steel between the end of the insulating sheath and the seating structure of the head, and often do not provide proper protection around the head itself where a locking head has been employed. The reliability and watertightness of the connection can be increased by welding pipes to the underside of the headplate, and arranging these telescopically over the insulating sheath of the anchor tendon. Alternatively, the end of the sheath may be extended into an expanded cavity, coaxial with the borehole (Fig. 18-17), in the anchored structure. The connection formed in this way must be sealed with a sleeve slipped over this connection, with

277

bonehole fi 150-110mm

Fig. 18-16. Prepared recess for the embedding of a VSL anchor head in concrete {Tarbela Dam) 1 — anchor, 2 — recess grouted after stressing the anchor, 3 — reinforcement

278

Fig. 18-17. Sealing of the connection of a Cona-Sol G. M. anchor head a) — anchor cased along its free length, b) — anchor fully cased 1 — trumpet, 2 — seal, 5 — secondary filling, 4 — smooth plastic duct, 5 — secondary grout pipe, 6 — tendon, 7 — corrugated plastic duct, 8 — spacer, 9 — primary filling, 10 — grout pipe

a rubber strap, or by some means compatible with the structural arrangement of the anchor. When the prestressing has been carried out, the head is covered with a cap made of sheet steel or reinforced plastic material to facilitate grouting of the head with cement mortar or synthetic resin under increased pressure (Fig. 18-18).

279

A very reliable anticorrosive protection was devised for the heads of the above-mentioned anchors installed at the pumping station for the water supply to Bratislava (Czechoslovakia). The steel casing pipe of the borehole, left in the ground as an external sheathing, was telescopically inserted into a steel pipe of larger diameter concreted into the anchored structure co-

Fig. 18-18. Protection of a 4.9 MN BBRV anchor head {Ringhals, Sweden) / — anchor head, 2 — grease

axially with the anchor (see Fig. 18-12). Such a telescopic connection of the casing pipe to the structure excludes the danger of a drop in the prestressing of the anchor together with a lack of response from the prestressed ground; such would be the case if there were a fixed connection between the casing pipe and the structure, because the casing pipe would then function as a compressed pile. The telescopically free connection is sealed with a rubber ring, and after prestressing of the anchor by grout. The internal insulating sheathing of the anchor, passing through the casing, is connected to a flange on the anchor head. The anchor has a cast-steel head in which the spliced end of the cable is concreted; this fixing arrangement is the least susceptible to corrosion. Considering that even where fixed heads are used the wires are pulled out to some extent from the head during prestressing of the anchor (Fig. 18-19), the connecting flange must have a watertight fixing to its seating surface and the insulating sheath of the anchor inserted into this flange must be sealed inside the latter with a flexible rubber collar. The head, including the spacer headplates on which the head is supported after prestressing, is covered by a protective wrapping which makes it possible to grout the head, the headplates, and the mouth of the borehole in the final stage of installation.

280

tensile force MN Fig. 18-19. Extraction of wires from an anchor head under tensile stress. (The wires are embedded in concrete within a steel cone-shaped bucket)

18.2.5

Anticorrosive protection of anchors of short service life

The previously described methods of protecting permanent anchors are i n many instances very laborious and costly. Their complex design often arises out of a lack of long-term experience (although the anchors in the Cheurfas Dam have served for 50 years), and a desire to overcome the initial mistrust of some of the investors in this up-to-date technology. For anchors of short service life, a simpler type of anticorrosive protection suffices: often just grouting the borehole in the root section closed off with a seal, and subsequently filling the remaining length of the borehole with cement or mortar following prestressing of the anchor. If the anchor tendon is to remain free, the borehole above the seal is not filled with grout, but the anchor wires, ropes or bars are provided with an anticorrosive coat. Temporary cable-type anchors can also be made with wires protected by a layer of zinc which is applied by dipping the wires in the molten metal, or by spraying them with zinc or aluminium (metallization). It must be emphasized that even these metals are attacked in the ground after a time, and therefore should only be used as a protection for temporary anchors. Such a metal coating should act as an anode, which then provides anticorrosive protection, even if it is locally damaged (pierced) by the galvanic effect. From experience, the service life of a zinc coat on iron in a moderately aggressive soil is about 10 years. Protective coats can also be formed by phosphate or chromate treatment, but as these coats are porous and are easily damaged mechanically, they are unsuitable for permanent anchors.

281

18.2.6

Anticorrosive filling of boreholes

As stated earlier, anchors passing through rock must be protected by an anticorrosive filling in the space between the borehole walls and the anchor, even if the anchors are already covered with an anticorrosive coat, wrapping, or pipe. This filling also contributes to the strength of the rock, and prevents rock decay by the action of air and water, etc. The best type of filling, and the method used to manipulate it, must be selected with these considerations in mind. Generally, the filling mixture is forced into the borehole under pressure, since this results in the cracks and cavities in the vicinity of the borehole being filled as well. In some cases, the filling of the borehole forms the only anticorrosive protection. The borehole walls are sealed with grout under high pressure, and soon after this grout has set, the borehole is rebored [106]. When the borehole has been flushed, the anchor cable, thoroughly cleaned along the fixing section, is inserted into the borehole. The fixing section is embedded in grout, and when this has hardened, the cable is prestressed. Following the prestressing, the free cable section, which is usually coated with a watersoluble oil for temporary protection (see Section 18.4), is rinsed with a detergent and embedded in grout. Instead of grout, bitumen compounds are sometimes used as anticorrosive fillings (Fig. 18-20). In the second reconstruction of the Cheurfas Dam in 1967 (see Section 24.3.1), the boreholes were filled with a heavy bitumen oil,

2

Fig. 18-20. Protection of 2 MN anchor heads used in the second reconstruction of the Cheurfas Dam 1 — anchor head, 2 — bitumen oil, 3 — load distribution plate, 4 — load distribution sill, 5 — sand, 6 — mortar, 7 — grouting pipe, 8 — levelling concrete course, 9 — base, 10 — sleeve

282

the specific weight of which was increased to 1,200 kg/m 3 by an addition of red lead [106]. The increased specific weight of the compound resulted in the expulsion of res : dual water from the otherwise sealed borehole (the borehole walls having been grouted before insertion of the cable). The initial sealing of the borehole walls was carried out to overcome the difficulties of inserting and insulating the cable in the presence of water, and also to create an additional protective zone around the anchors. Since the effect of this grouting was only to seal the cracks rather than make the hole watertight, the borehole walls were subsequently sealed by carrying out chemical injections in two stages. In the first stage the boreholes were injected with a dilute solution of water glass which had been left to gel for a considerable time; this solution was able to penetrate the rock to a depth of 30 to 50 cm around the hole. In the second stage, a thicker, rapidly setting solution was used. This prevented the thin solution from running back into the borehole under back-pressure from the rock. The insulating wrapping filling the borehole space around the cable had a viscosity of 400 poises at 20 °C, and 200 poises at 25 °C. Tt did not contain any aggressive additives which could threaten the anchor steel and it was electrically non-conductive. The insulating compound was diluted to obtain perfect envelopment of the anchor wires, and to facilitate the filling of the borehole. The anchors of the Muda Dam were protected in a similar way [215].

18.3 E L E C T R I C A L A N T I C O R R O S I V E P R O T E C T I O N

Electrical protection measures suppress simple soil-mediated corrosion, and neutralize the effects of stray currents. The effects of stray currents on an anchorage are usually insignificant, even in an area of strong electrical fields, because anchors are not large enough to transfer stray currents from one part of a field to another, as pipelines or cables may do. The various types of electrical protection function in the same way, namely they maintain the material of the anchor at a potential which either prevents corrosion from taking place at all, or only allows corrosion to occur at an acceptable rate. The potential required to protect steel (the so-called protecting potential) varies little over a wide range of conditions, and amounts to —0.85 V. For reasons of economy, electrical protection is mainly used for the passive protection of structures buried in soil [232]. Electrical anticorrosive protection includes not only the familiar cathodic protection but also various types of electrical grounding to divert and leak away stray currents; as far as anchorages are concerned, however, cathodic protection is the most important. This method of metal protection is of comparatively recent origin, and is used principally for underground pipe-

283

lines. Other underground installations are protected by this method only to a limited extent, although here too cathodic protection gives good results. The main reason for its underuse is the somewhat superficial knowledge of this relatively simple method generally held by technicians and engineers. Anchors fulfil the basic prerequisites for the successful application of cathodic protection, when the following conditions exist: a) An electrically conducting medium (electrolyte) occurs in the vicinity of the protected metal surface. b) The electrolyte enveloping the protected surface forms a sufficiently thick layer to carry a uniform distributed current towards the protected metal. c) The protected equipment does not have a complicated shape (large projections or hollows). The principle of cathodic protection rests in the induction of an electric current in such a way that the surface of the protected equipment is cathodically polarized. Thus a potential is set up which, on the basis of thermodynamic principles, limits or prevents corrosion on the cathode (Fig. 18-21). The

Fig. 18-21. Cathodic anchor protection by means of sacrificed anodes 1 — sacrificed anode (protector), 2 — insulating washer, 3 — anchor steel

Fig. 18-22. Relationship between current density and the effectiveness of cathodic protection

value of the protecting potential for steel is - 0 . 8 5 V, although in practice, it is usually maintained within the range - 0 . 9 to - 1 . 1 V. To achieve a maximum protective efficiency, an optimum current density has to be attained. Any increase over the optimum protecting current density results in practice in a reduced protective efficiency. The dependence of the efficiency of protection on the protective current density, for example in a bathing solution of 0.005 (M) ZnCl 2 (zinc chloride) is represented by the curve in Fig. 18-22. Some minimum current density values for the protection of metals are listed in Table 18-11.

284 TABLE 18-11 Minimum current density value for the protection of metals Metal

Medium

Minimum protecting current density [mA/m 2 ]

Author

Experimental conditions

Steel

0.001—0.1 % H2S04

600

Klement

Gentle mechanical stirring

Iron

0.0002—0.6 % NaCl

106

Bayer and Forel

Zinc

0.05 % KC1

1,500

Bulach

Steel

Highly corrosive soil with 0.5 % NaCl

400

Pritula

Iron

Sea water

170

Negrejev

Iron

Soil

16.6

Olsen

With damaged bitumen coat

Iron

Soil

0.7

Olsen

With undamaged bitumen coat

18.3.1

Gentle stirring

Circuit arrangement in cathodic protection schemes

The basic arrangement of a cathodic protection scheme is shown in Fig. 18-23. A protecting electrical circuit may be set up on this basis, the components of the circuit determining to a large extent what current will

W/w/X.

6^

Φ 7 ^

I ?>0

Φ

- +

source of current

3

Fig. 18-23. Cathodic anchor protection 1 — anchor, 2 — source of current, 3 — anode

Fig. 18-24. Electrical circuit for cathodic protection

285

be necessary from the source (Fig. 18-24). Rt to R6 denote the following, respectively: The resistance of the conductor connecting the source to the anode, the contact resistance of the anode, the resistance of the corrosive medium, the contact resistance of the corrosive medium on the protected metal surface, the surface resistance of the protected equipment, and the resistance of the conductor leading from the protected equipment back to the source. The total resistance of the cathodic protection circuit is given by the sum of the individual component resistances: R0 = Rl + ... + R6, and the total voltage drop in the protecting circuit is given by the sum of the voltage drops across each of the component resistances: E0 = Ei + ... + E6. The total output required depends on the necessary protecting current density. (Table 18-11) 18.3.2

Anodes and sources of electric current

One of the most important parts of the cathodic protection system is the anode, from which current is distributed over the protected surface. The placement of the anode should therefore be such as to distribute the current as uniformly as possible over the protected surface; in order to achieve this,. a system of anodes is often used (see Fig. 18-21). The Sigri Elektrographit Co. of Meitingen (GFR) has developed a system of electric protection, "Elprot", consisting of graphite anodes impregnated in a vacuum. These have a much longer service life than metal anodes. They are connected by specially designed cables to batteries in which graphite powder is used for the filling between the anodes. This system guarantees reliable long-term protection for structures embedded in the ground, using small amounts of power. The electric current for the system can be obtained from an external source (usually a mains rectifier), or can be created internally, by setting up a galvanic cell involving the protected surface itself and some less noble metal, which is then termed the sacrificed anode, or protector. In the latter case the system and its installation are considerably simplified, because output required from the source is very small. The protectors can continue to supply a low output for a long time, which means that this type of system is the more cost-effective; there are no operating costs, and no service or maintenance is required. External sources can nevertheless be used to advantage under certain conditions (large numbers of anchors, proximity of an electricity

286 supply, protection system already in operation for other installations nearby, etc. In Czechoslovakia, protectors are usually made from magnesium alloys, and are 10 cm in diameter and 80 cm long. Another type of protector is made from zinc and aluminium alloy. Below are listed some data pertaining to sacrificed anodes. (Table 18-111). TABLE 18-111 Characteristics of sacrificed anodes

Specific weight Theoretical current yield, [Ah/kg] Theoretical loss of weight, [kg/A year] Effective current per cent, of theoretical Actual current yield [Ah/kg] Actual loss of weight, [kg/A year] Potential relative to copper sulphate electrode [V] Part of potential more negative than potential of steel —0.85 V (relative to copper sulphate electrode)

Pure magnesium

Special alloy of magnesium H

Aluminium with 5 % zinc

1.73

1.94

2.92

2,200

2,200

2,870

39.5

39.5

29.5

49

55

39

1,080

1,210

1,120

80.5

71.8

75.7

—1.7

—1.55

—1.1

—0.85

—0.70

—0.25

Assuming an anchor length of 10 to 15 m and a diameter of about 80 mm, one protector of Czechoslovak manufacture will aiford protection to an anchor for at least ten years, if its insulation is poor or damaged. If the insulation is carefully made from fabric bandages, or with a layer of plastic material the lifetime extends to 50 or more years. Where anchors are positioned far apart, the sacrificed anodes are mounted directly on the anchors. In these cases, it is recommended that magnesium anodes with a variable resistor in the upper part and a fixed resistor in the lower part, be installed. It should be noted that to exceed a potential of — 1.1 V may harm the steel installation, causing steel brittleness in extreme cases, and a weakening of the steel-concrete bond. This danger, however, only arises when the current density needed for obtaining cathodic protection has been exceeded several-fold. The cathodic protection of a system of anchors is usually considered as a complementary part of the overall protection system, providing an extra line of defence in case of damage to the insulation. In this respect cathodic protection has the advantage of concentrating its effect on a limited area where some deficiency of the insulation has occurred, so that only small

287

currents are needed. In any case, cathodic protection gives confidence and reassurance to all concerned that their anchors will have a service life outlasting that of the structure which is being secured.

18.4 TEMPORARY ANTICORROSIVE PROTECTION

The material of an anchor is attacked by corrosion to varying degrees, depending on local conditions, and corrosion may set in as soon as material is delivered to the site, before the proper anticorrosive treatment is applied. Anchors which are to be protected from corrosion by grout, may be attacked before the space in the borehole around the anchor can be filled, either with cement grout or other protective grouting material. To overcome this problem, the surface of the wires, bars or ropes of which the anchors are composed, are provided with a coating guaranteeing anticorrosive protection for one month, or up to six months at most. Temporary anticorrosive coatings are made of paint-on materials based on oil with added inhibitors; resins or epoxytars are also used, but less frequently. All these materials are relatively expensive, and what is more important, they are difficult to remove, some of them presenting an obstacle to the formation of a cohesive bond between grout and steel. Thus for anchors or parts of anchors, the function of which depends on their adhesion to the grout-filling of the borehole, only those coatings can be used in which molecular interaction with the grout occurs. Given this requirement, the simplest coating is one of cement slurry, either pure, or modified with additions of various plastic materials. The problem with this kind of protection is that it may peel off, and drop down to clog the lower end of the borehole; this may seriously interfere with the later wrapping of the anchor with grout which is intended to provide permanent anticorrosive protection. Good results have been obtained with protective coatings consisting of water emulsions of special paint materials', their composition makes it possible to wash them off with water, either immediately before the anchor is inserted into the borehole, or after it has been embedded; in other types the physical properties of the coating are adjusted so that grout adhesion is not affected. An obvious requirement of these coatings is, of course, that they do not adversely affect the concrete. Examples of such protective coatings are, Rust-Ban 310 in the form of a water emulsion, and Rust-Ban 393, supplied as a solution. Both are used by the Dywidag Co. in the German Federal Republic. However, the makers of these coatings aim at developing temporary anticorrosive treatments which function efficiently until the final anticorrosive protection is applied, and which do not affect the adhesion of grout or other

288

materials used for the fixing of the anchor in the borehole. In Czechoslovakia, for example, the three-component reactive paint S 2008 is recommended; this creates a phosphate layer and an organic anticorrosive film on the surface of the steel. Wires for prestressing purposes treated with this paint and exposed to the atmosphere, have not been found to show any signs of corrosion for at least three months. The protective coating does not reduce the adhesion of grout to steel. The coats are applied by painting, spraying, or immersing in a bath of the protective material. Dipping in a bath is the most efficient application method. Coils of wire or rope which are to be used for the preparation of anchors should be protected under Polyvinylchloride or polyester wrapping during transport and storage. Coils of wire wrapped in this way together with a suitable desiccant, were found to be untouched by corrosion after a year of storage in an open, unprotected place.

Chapter 19 LONG-TERM OBSERVATION OF ANCHORS

The observation of an anchor is referred to as being long-term when the observation continues beyond 24 hours after the installation is carried out [120], The purpose of this observation is to record any changes taking place in the prestressing, or displacement of anchors as a result of temperature fluctuations, shocks, load variations of the anchored structure, changes in the state of stress of the rock, etc. The long-term observation of prestressed anchors provides important supplementary data to that obtained in short-term anchor tests, and it may also furnish valuable information concerning the anchored structure and the ground. In spite of the considerable importance of data obtained from long-term observation, there is still relatively little material available in this respect. In the following Section, the time-dependent losses of anchor prestressing which occur in the absence of external influence, will be discussed first. 19.1 LOSSES OF ANCHOR PRESTRESSING WITH TIME

Anchors which transmit a permanent tensile forca into the ground always show a drop in the initial prestressing with time. This loss is largely a result of the combined effects of relaxation of the anchor steel, and creep of the loaded ground. Relaxation is defined as a drop in the prestressing without deformation, while creep involves the deformation of a material under a permanent load. Knowledge concerning the relaxation of steel is now comprehensive, but much less is known about the creep of a rock or soil permanently under load from an anchor, there being few data available on the actual magnitude and distribution of stresses in the root zones of anchors. The losses attributable to creep in particular, can be considerable. 19.1.1 Relaxation of steel The characteristics of steel relaxation are wall knowa and are included among the data supplied by manufacturers from investigations on prestressed concrete. The relaxation losses in prestressed steel under long-term loading are usually within a range of 5 to 10 per cent.

290 On the basis of tests made on various types of steel [2], the losses caused by relaxation after 100 hours of loading are found to be approximately double the losses occurring after 1 hour of loading, 80 per cent, of the loss of stress after 1,000 hours of loading, and 40 per cent, of the loss after 30 years of loading. The relaxation values vary in relation to the loading of the steel. When steel is stressed to 50 per cent, of its strength, the relaxation losses are negligible, but they rapidly increase with higher loads, and also increase significantly with temperature above 20 °C. The introduction of stabilized wires and strands has reduced stress losses from the 5 — 10 per cent range of ordinary stress-relieved steel to 1.5 per cent. at 75 per cent, of the guaranteed tensile strength at 20 °C. In steel loaded for long periods, losses of prestressing due to creep deformation have also been found. These losses are negligible, however, compared with the effects of relaxation. 19.1.2

Creep of the ground

Creep of the ground under load arises from plastic compression, or failure of the rock or soil under the stresses brought about in the zone affected by the load. In the case of prestressed anchors, creep occurs primarily in places of concentrated stress —near the anchor root and below the anchor head at the surface of the anchored structure. 19.1.2.1

Behaviour of hard rocks

In dense hard rock in which the loading stress is accommodated by the strength of the rock, some additional compression of the natural planes of separation (joints, cracks) takes place. The fall in anchor prestressing rapidly diminishes with time, until an equilibrium state is reached. In the case of bar anchors, the total loss is approximately 20 per cent, of the original prestressing, but in cable anchors of greater ductility the loss is considerably less, particularly in longer anchors (more than 10 m). The original prestressing can easily be restored by increasing the anchor tension using the stressing equipment; any subsequent drop is substantially less. Where strong rock is concerned, additional stressing is usually carried out twice: after 24 hours, and then after a week or two. The time-dependent development of prestressing losses in two different bolts and long bar anchors is shown in Fig. 19-1. Creep in strong, hard rock is very small, even under high and prolonged loading. According to the PCI [238] the prestressing losses of rock anchors reach 3 per cent, after 7 days and are attributable entirely to the relaxation of the steel. Long-term observations of anchored dams also show that the

291

20

30 time

40

50

[days]

5Ö0, 300

20θΗ 100 0

- — - 00
time

(months)

^^: Fig. 19-1. Drop in prestressing in different types of anchor with time A — bolts; 1 — wedge bolt 1.80 m long (the ultimate strength of the bolt not exceeded), 2 — wedge bolt (ultimate strength exceeded). (Both bolts fixed in strong migmatite). 3 — bolt in cement mortar (60 cm root) in clayey shale [227], B — bar anchor 11.00 m long, 5.00 m root, C — bar anchor 14.00 m long, 5.00 m root [159]

losses reach a maximum of 10 per cent, and are caused more by the relaxation of steel and creep of the concrete than by creep in the bedrock. The longest monitoring of anchor prestressing has been carried out at the Cheurfas Dam in Algeria. After three years the losses were 4 per cent., and after 18 years they had reached only 5.5 per cent [120]. The anchors, prestressed to 10 MN, were fixed in strong sandstone (see Chapter 24). Comte [34] recorded losses of 4 —8 per cent in 1,250 kN BBRV anchors fixed in very variable fissured argillaceous schist in the Nendaz Cavern. These losses were notably within the 10 per cent, margin allowed, and the

292 greater part of the loss was found to occur in the very early stages of a fiveyear period of observation. In the course of prestressing two test anchors (fixed anchor length 6 m, diameter 99 mm), Barron et al. [8] subjected one of the anchors to three loading cycles prior to lock-off, whereas the other anchor was loaded directly with the lock-off load. Both were installed in jointed granite (Fig. 19-2). The tension in the first anchor remained stable throughout the no 135 130 -^125

\l20 \l15 |Wj ^105

wo r 95

90

r*> ~ 10 V \ -0)

_—-* v

*-*-Xs

10 3 4 * 6 time(monfns)

1

8

9

10

Fig. 19-2. Comparison of anchor prestressing performance with time and temperature [8] 1 — length of tendon 59.5 m, anchor not repeatedly loaded before lock-off, initial load 133.8 kN, 2 — length of tendon 10.1 m, anchor subjected to three loading cycles before lock-off, initial load 119.5 kN observation period, whereas a stable state was achieved in the second anchor only after a marked loss had occurred (by the end of the first week) owing to the closing of fissures in the rock. The authors concluded that it is better to raise the loading to its maximum value through a series of cycles, as a means of minimizing tension losses after lock-off. There also appeared to be a temperature effect on the apparent tension in the shorter anchor. Möschler and Matt [138] presented data on the performance of a 1,330 kN VSL anchor (root length 4.50 m) after test-loading it to 1,725 kN in fractured calcareous schist in Waldeck Cavern (Fig. 19-3). 19.1.2.2

Creep of soils and soft rocks

In soft rocks and soils deformation arising from ground compression is considerable, and the attenuation of this deformation is relatively slower. Very marked and long lasting deformations have been found in cohesive

293

clayey soils and in fine, uniformly grained sands [154]. In these soils large creep displacements of the anchor root, and plastic flow of the soil around the root, take place at the ultimate load. The displacements continue to increase with time and the required tension in the anchor cannot be maintained permanently; thus the load-bearing capacity drops and the danger increases that the anchor root will be torn out of the soil. 1.33.Λ

\j\\ . / Ρ^ Ν Λ

.132

2 3

1.131

I

1.29

\ f

1000 hours i i

1M

-c:

■c:

5^

<=5 I

I

1.23 1.27

CO

*

6 8 time (months)

10

11

1t

Fig. 19-3. Performance with time of a monitored anchor [138] 1 — initial reading, 2 — design load (1.33 MN), 3 — theoretical tendon relaxation curve, 4 — actual anchor performance, 5 — lowest stress recorded (loss of 0.04 MN = 3 %)

It is therefore important to know how creep deformation develops with time, especially where anchors are to be installed in the more compressible types of ground. Generally, the creep-time relationship for permanently loaded anchors is near-exponential. The results of anchor tests carried out by Ostermayer [153] in uniformly grained sand are shown in Fig. 19-4. The gradients of straight lines in the time-displacement diagram give the value of the coefficient of creep, ks, which increases with the load tension in the anchor. That part of the displacement attributable to the partial separation of the tendon from contact with the grout and to relaxation, has a ks value of approximately 0.4 mm. Values of the coefficient above this signify creep along the root/soil interface. It is possible to calculate theoretically the expected long-term creep displacements of the root, using the creep coefficient found in an anchor test, and thus form an idea of the losses of stress that can be expected with time. Ostermayer recommends that the admissible limit of the coefficient of creep should be set at 1 mm under a load of 1.5PW, for the testing of permanent anchors in cohesive soils. A value of 1 mm for ks theoretically corresponds to a displacement of 6 mm occurring in a time interval stretching from 30 minutes to 50 years. The results of many tests show that anchors 20 —25 m long with long

294 time 10

/minutes) 100

1000

Fig. 19-4. Time-displacement curves and creep coefficient for different loads in a uniform sand. Two anchors monitored. Values obtained in the first anchor are marked with circles (o), in the second anchor with triangles (Δ) [154]

580kN

Δ

\580kN

2610 20 30 W 50 60

80 90 100 110 120130 W1S0 duration of prestress[days]

Fig. 19-5. Extraction distance of an anchor under a sustained force of 345 kN. The anchor was fixed in loess loam at a depth of 11 m by a bulb 35 cm in diameter

grouted roots of 10—15 cm diameter register a prestressing loss of about 6 per cent in hard clays, and 12 per cent, in stiff clays, because of creep [154]. It is interesting to note that these losses are usually registered within the

295

first 2 —4 months following the prestressing of the anchor and do not increase thereafter. Measured values are generally lower than those calculated using the creep coefficient obtained during the initial loading. The test result of a cable anchor fixed at a depth of 11 m in loess loam with a bulb concreted at the foot of the blasted-out borehole (permanent loading force 345 kN), is shown in Fig. 19-5. Losses of prestressing caused by creep in soils and soft rocks may be compensated for by repeated additional stressing at increasingly longer time intervals (up to one year). This can be done provided that there is no danger of the load-carrying capacity of the root being exceeded, and that the anchor was designed for additional prestressing. Anchors cannot be installed in highly compressible soils with large amounts of organic matter, or in very soft ground (made-up ground, loose sand) of low consistency ( < 0.9) or high liquid limit ( > 50 %), because of the large creep deformations that would occur. 19.1.3

Observation of prestressing losses caused by relaxation and creep in production anchors

The monitoring of prestressing losses caused by relaxation and creep is usually prescribed or recommended for a proportion of all production anchors. The FIP [120] recommends that 10 per cent, of anchors should be monitored. The French Standard prescribes the monitoring of 5 to 15 per cent, of permanent anchors (depending on the total number) for at least 10 years. In the first year the anchors are inspected every 3 months, in the second year every 6 months, and then at intervals of one year. In Great Britain and the South African Republic prestressing loss is measured after 24 and/or 48 hours following the prestressing of all temporary and permanent anchors. If the results are satisfactory, observation is continued on 5 per cent, of all production anchors for one year [120]. The admissible change in the prestressing of an anchor is usually 0.1PW. In Germany and in other countries the magnitude and development of tendon displacement under constant load are followed in detailed acceptance tests of production anchors. An anchor is satisfactory if the displacement increases in proportion with the logarithm of the time, or if displacements decrease with time. Long-term observation of prestressing losses, and their rectification by additional prestressing of the anchor, are possible only in those anchors which have permanently free tendons (between the root and the head of the anchor).

296 19.2 CHANGES IN ANCHOR PRESTRESSING DUE TO EXTERNAL FACTORS

Various external factors can bring about changes in the loading of an anchor, leading to a permanent reduction of the prestressing in production anchors. This may be caused, for example, by shocks in the anchoring medium, or by a variable or fluctuating load exerted by the anchored structure. Other effects may even result in an increase in anchor stress, e.g. changes in temperature, changes in the equilibrium stress system of the ground, etc. Such changes in anchor prestressing can markedly affect or even impair the function which the anchor was intended to fulfil. 19.2.1

Shocks occurring in the anchoring medium

The highest intensity shocks recorded in anchor-holding rocks are most often the result of blasting. Shocks can also be caused by heavy machinery, and by earthquakes in seismically active regions (see Chapter 8). Shocks are the cause of prestressing losses in anchors much greater than those caused by long-term static loading; in extreme cases (frequent occurence, high intensity) shocks may lead not only to prestressing loss, but also to a substantial reduction of the load-bearing capacity of the anchor. As in the case of time-dependent losses, ther is currently little data available on the effects of shocks on anchors. In the USA, research has been carried out on the effects of blasting on bolt anchorages in horizontally stratified dolomite in mines [68]. There was a marked drop in prestressing when explosives were used within 3 m of the anchors. This drop was approximately 36 times greater than the drop in prestressing that occurred in the same bolts over a similar time interval under static load (Fig. 19-6). At a distance of more than 5 m the effect of ordinary blasts was insignificant. Conditions for the propagation of seismic waves are particularly favourable in hard and compact rocks, and in rocks fissured along cleavage planes. Shocks may bring about a change in the prestressing and load-bearing capacity of anchors in poorly compacted non-cohesive soils, and they can have particularly unfavourable consequences in cohesive soils with labile thixotropic properties. Careful laboratory tests on soils, and in-situ testing of anchors, are essential before construction work is started. Of the various types of anchors, bolts with a mechanical fixing in the borehole suffer the most from shocks; cemented or combined anchor fixings (particularly those with synthetic resins) are much less sensitive. Anchors fixed in the rock or soil by means of an expanded root (abutting base) [214] show a higher resistance to shocks than anchors fixed by means of a long root. A mechanical device has been developed in S. Africa [152], which

297

allows a gliding movement of the tendon in the base, and gives very efficient shock protection to prestressed bolts with mechanical bases. Whereas normal bolts were found to lose all their load-bearing capacity, and collapsed along with the rock after a blast was let off in the anchored roof, yielding bolts held the rock roof intact and only required tightening. Some types of yielding bolts are described in Section 13.1.5.

40Ö L-Li

2

1

3

1

1 I i

U 5

6

U

7

β

i

9

i

i

i

i | i

i

i

i

i

10 11 12 13 1k 15 16 17 ^ - time[days]

Fig. 19-6. Drop in prestressing of bolts with time and as a result of shocks from blasts [68] a> byc,d — tests bolts, 1 to 30 — successive blasts

Restressing of anchors within range of shocks must be carried out regularly, if the anchorage is to maintain its action on the structure or ground. Long cable anchors are less affected by shocks than short bolts. In Czechoslovakia, several measurements have been carried out on anchored structures to ascertain the amplitude of variations in the anchoring forces during blasting operations. The variations were found to be very small. This may be explained by the fact that any variation in the anchoring force must be accompanied by a simultaneous change in the distance between the anchor head and the root. In blasting operations, a vibratory motion of the bedrock takes place within a frequency spectrum of 5 to 50 Hz. The amount of the charge is arranged so that the velocity of the vibratory motion does not exceed the limit for the structure that is threatened. A velocity, Fmax = = 80 mm/s, is taken as the highest permissible limit. As an example, let us suppose that the bedrock suffers a harmonic vibration so that the maximum displacement of the ground surface is 2.5 mm at a frequency of 5 Hz, and only 0.25 mm at 50 Hz. If the free length, L, of the anchor is 10 m, and this is prestressed so that the relative elongation, ε, is 0.006, the elastic elongation, Δ/, will be 60 mm. Thus the amplitude of variation of the anchoring force

298

at a frequency of 5 Hz is 4 % of the total anchoring force, and only 0.4 % at 50 Hz. Considering that blasting operations give rise only to a short-term change in the loading, long cable anchors are not seriously threatened by a large drop in the anchoring force caused by blast shocks. 19.2.2

Variable loading of anchors

Continuing long-term rapid variations in the anchor load can have an adverse effect on the maintenance of tension in the tendon, and/or fixing strength of the root. The effecs of a rapidly fluctuating load on an anchored structure has been studied in Czechoslovakia. The structure in question was a concrete compensator block; the weight of the rotor was 200,000 kg, and the rotation speed was 12.5 Hz. The block was built on gravel and sand 10 m thick covering strata of clayey shales and sandstones in which the anchor bulbs were fixed. After 10 years of service, a drop of only 17 % in the prestressing of the anchor was observed. At more extreme fluctuations of the load (for example, those occurring in the anchored concrete blocks of high masts involving load variations of up to 50 %), a large drop in the prestressing is to be expected. In these cases, regular checking of the anchoring force must be provided for, and a facility for additional prestressing must be included. 19.2.3

Changes in temperature and the state of stress of the anchoring medium

Changes in temperature and changes in the stress state of the anchoring medium can bring about a decrease or an increase in the anchor stress, and unless this is foreseen in the anchor design, these effects present a real danger to the anchor, and can eventually be the cause of failure. Temperature changes bring about expansion or contraction in anchors and in anchored structures, these dimensional changes depending on the coefficient of thermal extensibility for the material concerned (i.e. the effect of the same temperature change is different with respect to the anchor and the anchored structure). The effect on anchors of changes in air temperature is negligible, since anchors are for the most part below ground, and the steel of which anchors are made is of high ductility. Anchored structures, on the other hand, with large areas exposed to the atmosphere and direct radiation from the sun, can expand or contract and thus affect the prestressing of the anchor to some extent. Since these effects operate slowly and are restricted to within a narrow range, they do not threaten the function of anchors in most cases. Much greater effects on anchor prestressing can be caused by changes in the

299

state of stress of the anchored structure, especially where such changes could not, or were not considered as part of the static analysis. A very dangerous phenomenon is the increasing tension that occurs in short bar anchors supporting underground excavations when the latter are extended, or when new excavations are going on nearby. The increase of stress within the rock can give rise to a load exceeding the tensile strength of the bolt. In such cases a minute displacement of the anchor head suffices to decrease the load to within the admissible limit, so that the bolt can maintain its stabilizing function. As in the case in dealing with the effects of shocks, the problem is solved by using bolts with a small automatic yielding capacity, which comes into operation as soon as the tensile force reaches a critical value. Another solution is to supplement the reinforcement with further bolts. Sometimes a foreseeable increase in the load is allowed for by setting a lower value for the initial prestressing of the anchor, or by installing the anchor without any prestressing. The most reliable assessment, of course, is provided by a systematic observation of the changes of stress (deformation) occurring in the rock, and measurement of changes in anchor prestressing. Thus appropriate steps can be taken in time, these usually involving a strengthening of the reinforcement. Sometimes special measuring anchors are installed solely for the purpose of observing the changes which take place in the state of stress of the ground with time, and under the influence of external factors. Of the latter, the effect of further excavations on the surface or underground is of primary importance. Changes in the prestressing of stabilizing anchors caused by a change in the state of stress and of pressure within the rock or soil medium in the surroundings of the excavation, are often observed on the supports of underground openings or on the sheeting walls used in the course of construction operations (see Figs. 20-16 and 22-30). The measured values are compared with those considered in the design. From the results the correctness of the calculations in the design, and the effectiveness of the support, can be assessed.

19.3 INSTRUMENTS FOR THE MEASUREMENT OF ANCHOR PRESTRESSING

Long-term and short-term observations of changes in the loading of prestressed anchors are carried out with numerous types of instruments (load cells), working on mechanical, hydraulic, electrical, vibrational and photoclastic principles. According to the system used, these instruments have different ranges of measurement, different degrees of precision and independence from external influences, and different susceptibilities to damage. They

300

are usually slipped on to the anchor tendon below the anchoring or stressing head, and care must always be taken to ensure that they are loaded centrically (evenly around the perimeter of the instrument). Their condition must be checked regularly. 19.3.1

Mechanical instruments

These operate on the basis of elastic deformations taking place in various types of steel washer or steel spring. Their measurement range and response are small, but they are very robust. A simple monitoring of the stress in short bar anchors can be achieved with the aid of calibrated spring washers placed beneath the tightening nut. The degree of compression of these washers indicates the change in stress, and can be measured. In ore mines, steel cup springs are sometimes used; these become inverted with respect to their concavity as they are forced against the nut (Fig. 19-7) under a critical pressure from the rock. This condition indicates that the maximum prestressing of the bolt has been reached.

Fig. 19-7. Dish-shaped washer made by the Elbroc, Bateman Co. A — before loading, B — after loading the bolt to more than 350 kN

Fig. 19-8. Resplat spring washer (Bachy) 1 — removable cover, 2 — concrete, 3 — anchor head, 4 — sensing element, 5 — adjustment distance of the setting of the sensor, 6 — spring washers, 7 — anchor

Further progress in this direction is represented in the Roof bolts' 'lockplate'\ which functions as a nut at the same time (see Fig. 16-8). This washer is slipped on to the unthreaded bolt rod, and after prestressing, the washer grips the rod strongly and reliably by means of self-gripping. The curved and triangular shape of this washer guarantees support at three points on the rock

301

surface. The deformation of the lockplate is so consistent that its various stages of deformity can be used as a visible guide as to the different tensile force in the bolt and stresses or movements of the bolted rock formations as illustrated in the photographs. Bolts with these heads can be restressed, if necessary. The French Bachy Company makes use of a set of spring washers to check changes in the prestressing of anchors with large loads (up to 10 MPa). The washers are covered, together with the anchoring head, by a wrapping containing sensing elements which automatically register anchor deformation beyond an admissible set value (Fig. 19-8). 19.3.2

Hydraulic instruments

These instruments essentially consist of a closed pressure vessel which is filled with oil and is connected to a manometer. Their advantages are: direct reading of pressures on the manometer scale, small dimensions and weight of the instrument, and considerable resistance to damage (with the exception of the manometer). Hydraulic instruments can be made relatively easily by constructing a small pressure vessel with an outlet for the manometer (Fig. 19-9).

Fig. 19-9. Hydraulic dynamometer BE-5 of Ostroj (Czechoslovakia)

Precise hydraulic load-measuring instruments for anchors are made by the German firm of F. Gloetzl (Fig. 19-10), for loads of up to min. 250 and/or max. 5,000 kN. These instruments weigh from 4 to 125 kg. They are accurate to ± 1 %, and the thermal error is only 1.2% of the measured range at a temperature diiference of 20 °C. The manometer calibrated for direct measurement can be provided with contacts for a signal light which is switched on when the set limit force has been reached.

302 Fig. 19-10. Hydraulic instrument for measuring anchor load (Maihak) A — view of complete instrument, B — Cross-section of instrument; dimensions A, B, C, D, E vary according to the capacity of the instrument, 1 — anchor, 2 — pressure vessel containing liquid, 3 — equalizing washer, 4 — pressure gauge

B)

19.3.3

Photoelastic instruments

In these instruments, the deformation of an optically sensitive material takes place. A sensitive glass disc is fixed in a strong steel body which is slipped on to the anchor tendon below the head (Fig. 19-11). The stress is measured on a portable optical gauge equipped with a red filter and an internal light source. The pattern on the sensitive material is compared with the standard patterns of force lines, and the bolt tension is then ascertained from a conversion diagram. The accuracy of the reading is within ± 1 %. Such instruments are made, for example, by the English firm of Horstmann, with measuring ranges of 0 - 20 kN up to 0 — 6,000 kN. They are relatively cheap, simple to use and are unaffected by external influences. 19.3.4

Electrical resistance instruments

These enable measurements to be made remotely on a portable apparatus which registers changes in the electrical resistance of loaded elastic measuring elements fixed in a strong steel hollow cylinder. The instruments are generally

303 Fig. 19-11. Horstman optical load meter in service a) — small type (up to 100 kN) on the bolt head, b) — large type (up to 400 kN), installed in cast-steel tubing forming tunnel reinforcement (in the left bottom corner the portable optical gauge is shown in use)

0

affected by external humidity. The reading apparatus is simple, being essentially a voltmeter. The accuracy is at short-term test + 1 °/ 00 a t long-term test no greater than ± 1 % of the measured range. These instruments are made in various sizes by several specialized firms, such as Terrametrics in the USA (Fig. 19-12) or the Swiss firm of Huggenberger (Fig. 19-13). Instruments made by the latter firm have a compensation facility for balancing a non-uniform loading around the cell perimeter.

304

Fig. 19-13. Huggenberger A. G. load meter. The dynamometer is shown on a bar anchor in (A) and the reading equipment is shown in (B)

19.3.5

String instruments

These are among the most reliable and most accurate load-measuring instruments. The measuring system is based on the vibration of three or six strings fixed in a cylindrical body with a central opening for the anchor. The vibration of the strings, induced by a vibration exciter in the reading apparatus, changes with the load. These instruments are made by the renowned German firm of Maihak, with ranges of 0 to 200, 0 to 500, 0 to 1,000, 0 to 2,000, and 0 to 3,000 kN (Fig. 19-14). They can be read directly or from a remote point, and are equipped with automatic recording. 19.3.6

Tensiometric instruments

Satisfactory measurements of the load in prestressed anchors can be obtained by using the well known strain gauge strips; these are fixed on the walls of a loaded steel cylinder, the deformation of which is then registered

305 Fig. 19-14. Maihak string dynamometer

electrically. An instrument of this type with a range of 1,000 — 3,000 kN was made at the Research Institute of Civil Engineering (Czechoslovakia); it is shown in Fig. 19-15 together with the reading apparatus. The accuracy of measurement was + 1 %. The load meters made by the Swiss Stump Bohr AG and Proceq SA (Fig. 19-16) are based on the same principle.

Fig. 19-15. VUIS tensionmetric dynamometer in use on a dam site (Czechoslovakia)

306

ill

Fig. 19-16. Strain gauge dynamometers DMS of Proceq SA for compressive forces of 1120, 2020, 3030, and 5050 kN

Chapter 20 A N C H O R I N G OF U N D E R G R O U N D

EXCAVATIONS

The anchoring of underground excavations was one of the first ways in which anchoring technology was applied, and is now the most widely used form of anchoring. The first reports on the strengthening of rock with steel bars date from before 1890. The reports relate to reinforcement work in the coal mines of North Wales, and work carried out in the USA before 1905 [61]. In Central Europe, bar anchors fixed in rock were used for the first time in 1918 [208] to secure the roof of an underground excavation in the Mir mine in Upper Silesia (now in Poland). In Czechoslovakia, the first successful use of anchored reinforcement in mining was made as early as in 1926, when shales were secured in the wall of a dipping shaft. The use of anchors, however, did not become widespread at that time. The general use of bolt reinforcement as a replacement for timbering in mines began in the USA during the Second World War, and then spread all over the world as further developments took place in the drilling technology used in mines. This made for quick and cheap drilling of anchor boreholes, and brought with it a better theoretical knowledge of the mechanics of rock masses. In the mines of the USA, more than two million bolts were being fixed per month by the end of 1952. In the fifties, this method was used on a large scale for the first time in the excavation of tunnels and underground caverns [172, 137]. It became widely used in underground constructions, not only as a temporary reinforcement instead of timbering and other types of frame structure, but also as the main permanent support system for the rock faces of excavations, instead of masonry and concrete linings. It also gave birth to a new tunnelling technology—the New Austrian tunnel-driving method.

20.1 PRESSURES ACTING ON ROCK SPACES, AND THE CALCULATION OF A N C H O R A G E PARAMETERS

An underground excavation disturbs the equilibrium stress state of the rock mass, and as a consequence, lumps of rock become loosened and fall from the exposed rock face, the rock is forced towards the excavated space, and the support system that is constructed experiences an additional stress load. These effects can generally be described as manifestations of rock

309

pressure. The source of this pressure is principally the force of gravity, although sometimes the residual stresses of orogenic activity within the earth's crust, including forces responsible for the formation of the surface relief, are also operative. Usually only the weight of the overlying rock above the excavation is considered. For a given depth below the surface, a vertical stress, σν = y . K acts within the rock mass, as well as a horizontal stress, σΛ

1 - v

σν = Κ.σν

(Fig. 20-1)

where h = the depth below the surface, y = the volume weight of the rock, v = the Poisson number of the rock.

£.-*"

\r

Fig. 20-1. Theoretical stress pattern in the vicinity of a circular opening in strong, homogeneous, isotropic rock with Poisson number v = 0.2 (according to K. Terzaghi), σν — vertical stress, ah — horizontal stress, σ0 — stress at zero level, as — stress at level S after completion of opening breaking, H — depth below surface

If the slow changes taking place in the stress pattern of the rock during excavation are considered in terms of Mohr's representation for homogenous medium (Fig. 20-2), then circle 2 denotes the state prior to excavation. Circle 3, which touches the rock failure curve, represents the limit state of the rock load-bearing capacity at the commencement of excavation, and circle 4 represents the substantially increased tangential stress developing in the rock as excavation proceeds (the radial stress being zero). The rock is usually unable to withstand the increased tangential stress, and consequently fails (loosens). Failure occurs initially in the roof, where tensile stresses occur (as shown in Fig. 20-1); the rock is unable to withstand these to the same extent as compressive stress. Only when the weight of the overburden

310 Fig. 20-2. Stress values in a rock mass before and after the breaking of an opening, represented in a Mohr's diagram (ace. to A. Hugon) 1 — envelope of full rock strength, 2 — circle representing stress state in the mass before the breaking of the opening, 3 — stress state at the moment of rock failure, 4 — stress state in the natural arch after completion of the opening (radial stress = 0), p — radial stress needed to secure the face against collapse

increases substantially, do the lateral stresses at the sides of the excavation grow to such an extent that disintegration and loosening are inevitable there also. If collapse at the sides is to be prevented, radial stresses (minimum pressure, p, see Fig. 20-2) must be created in the excavation face to bring the Mohr circle circumscribed over the difference of stresses below the curve of failure (circle 3). Such a pressure would have to be very considerable. In most cases however, it is not necessary. After some loosening of the rock at the excavation face, the stress increases and moves deeper into the rock mass where it is easily accommodated by the rock (Fig. 20-3). The loosened rock nearest the exposed surface must be supported or strengthened to prevent collapse under its own weight. If collapse occurs, the excavation has to be expanded and thus the stress is again increased, with the result that the stress is also shifted even further into the rock. Traditional timber, steel, or concrete supports for underground excavations support the loosened rock and prevent it pressing inward into the excavated

Fig. 20-3. Natural arch zone created in the rock when the stress has shifted inward from the rock face /—zone of reduced stress around the opening, / / — zone of increased tangential stress (of the natural arch), σν, ah — vertical and horizontal stresses within the mass before excavation, H0 — depth of the centre of the opening below the surface

311

space. However, the construction of such reinforcement is usually very time-consuming, and the irregular points of contact with the broken rock face of the excavation are unsatisfactory, since high concentrations of pressure are created both in the supported rock and in the supports. This leads to greater loosening of the rock in the vicinity of the excavation, and increased pressures. Anchors function on a different principle: they either connect the loosened parts of a strong rock with more stable regions of the mass, or they strengthen these parts by reinforcement and prestressing, the prestressing of the anchors being especially responsible for transmitting a definite load (see Fig. 6-1). Another advantage is that the anchorage can rapidly be put into action before the unsupported rock has been loosened too much. Both types of support in their usual form, however, are only able to withstand the full weight of the rock overlying the excavation to relatively small depths (approximately 10 m). Larger loads would inevitably lead to destruction. Over a certain depth below the ground surface collapse cannot occur, as the load-carrying function of the excavated rock is taken over, as mentioned above, by the surrounding rock mass; the reinforcement only supports the weight of the loosened rock in the immediate surrounds of the excavation. The transfer of the load-carrying function of the excavated rock to the surrounding rock can occur in either of two ways: either by the formation of a natural arch in the rock mass (the arch being supported by the undamaged rock at the sides of the opening), or by the formation of a rock beam resting at either end on the surrounding rock mass. The details of both theories follow from a consideration of the pressure zone in the rock mass lying over the excavation. 20.1.1 Rock beam theory This theory is applicable particularly to rectangular excavations in rock with strong, approximately horizontal bedding planes. Beds of smaller load-bearing capacity can be supported by anchors fixed into the more competent overlying rock, or a rock beam can be formed by extensive anchoring of such beds. In the first case, in which the roof is suspended from a load-bearing bed, the bolts anchored into the strong rock are tensioned by the weight of the suspended rock, and their length is determined both by the distance of the load-bearing bed from the roof of the excavation, and by the length required to fix the bolt into this bed. Other parameters of the anchorage can be determined as follows: The spacing of bolts, / r , is given by

312

WTT where Fs xt hs y

= = = =


·

area of cross-section of the bolt, excluding the thread (cm 2 ), permissible tensile stress of the bolt material (10" 1 MPa), thickness of the suspended strata (cm), volume weight of rock (kg/cm 3 ).

The bolts are always prestressed to the assumed value of the load, P (given by P = y . hs. / r 2 ), in order to verify their load-bearing capacity and reestablish (at least partially) the stress state in the rock existing before the commencement of excavation. The extreme bolts are placed as close as possible to the walls of the excavation, because the support zone of the rock beam can only be assumed to extend for a short distance back from the face. This distance is theoretically given by h I tg 45 —— J, as with a natural arch (see Fig. 20-6). The distance lr is sometimes assessed by considering the admissible tensile stress on the lowermost bed of the supported strata, between the two bolts. The bed is assumed to be partially fixed in the anchoring point. The permissible span is: / 2σ0 .hf h = J—-—V m .q where σ 0 ht m q

= = = =

(cm),

the tensile strength of the rock under the influence of bending, the thickness of the lowermost roof bed, safety factor (usually 2), uniform load of the bed arising from its dead weight.

If l'r < / r , either the spacing of the bolts must be equal to l'r and narrower bolts used, or the bolts must be connected by flangeplates in one direction, whilst the spacing is reduced to l'r in the other. It is also possible to secure the roof bed with wire netting, fixed to the roof with bolts spaced at the distance lr apart. The latter procedure is recommended as the most reliable means of preventing the loosening of the rock between bolts; the tensile strength of rock under the influence of bending is highly variable because of the presence of many transverse cracks and fissures, and is therefore preferably not taken into account. The fixing depth of the bolt in the load-bearing rock is determined by the requirement that the resistance to extraction of the bolt be greater than the bolt strength. It has been shown in tests that a dish-shaped body of rock is torn out by the anchor from a strong compact rock mass, whereas in fissured rock the natural planes of separation have a strong influence on the shape

313

of the torn-out part of the rock, which in this case closely resembles a regular cone with its apex (apex angle 90°) at the bolt fixing point (Fig. 20-4). From the condition: Fs. σ] ^ π . hu. Th

/

/

sin 45° '

\

a /

_l_y\4f

*-A

\

er 1

• ^l^

Fig. 20-4. Extraction of a bolt from load-bearing strong rock 1 — probable surface of separation in compact rock, 2 — surface of separation assumed in the analysis, 3 — roof, hu — fixing depth in load-bearing bed

the fixing depth can be derived: _

10.22 .F..o>

where Fs = area of cross-section of the bolt, ast = tensile strength of steel, xh = shear strength of the rock. If the bolt is fixed in the load-bearing bed with cement, the required fixing length /„ is usually greater than AM, so that the total bolt length in the rock will be:

/= K + K

or

/ = K + /.

If the level roof of an opening is formed of strong, thinly bedded rock, the individual beds can be held together by bolts to form a single beam. The strength of beds locked in this way is considerably greater than that of unconnected beds. L. A. Panek [162] has made a theoretical study of the strengthening of a plane roof in this way. He compiled a nomogram from which the parameters of the bolt anchorage can be derived; it was assumed that the roof beds are of identical thickness and strength, and that the bed surfaces show no resistance to reciprocal movement. This nomogram has been converted to the metric system by R. Kvapil and K. Luffr, who also extended it to give some of the characteristics of the rock, as determined by the strength and thickness of the beds (Fig. 20-5). Information can be readily

314

derived from the nomogram as to the effect of the proposed bolt anchorage under different conditions. A coefficient of consolidation of 1.5 to 2 may be considered satisfactory, since this will mean that the downward displacement of the roof is reduced by 33 to 50 per cent, at the collapse limit, compared with a roof without anchorage.

Fig. 20-5. Nomogram for the design of bolt anchorage in stratified rock over openings (according to L. A. Panek, modified by R. Kvapil and K. Luffr)

20.1.2

Natural arch theory

This theory is applicable to all types of compact and fissured rocks, to soft rocks, and even to soils. When the rock beam which was considered in the previous section fails, a natural arch develops over the opening, even in a stratified rock mass. The natural arch corresponds to a zone of increased stress which exists within the rock mass and is unaffected directly by the excavation. The pressure of the overburden is supported by this natural arch and transferred on to the sides of the opening and hence into the substrata. The weight of the loosened rock underneath the natural arch may load the support of the opening (see Fig. 20-3). It is therefore essential to know the position of the natural arch, at least approximately, and to calculate the pressures that will arise, as far as the support is concerned.

315

In subterranean constructions, the procedure of M. M. Protodjakonov [168] which is based on the theory of unconsolidated materials, is generally used. The lower edge of the natural arch directly over the opening is represented by a parabola (Fig. 20-6). The maximum thickness of the rock surface of ground

Fig. 20-6. Compressive zone of parabolic shape over an opening (according to M. M. Protodjakonov)

threatened by collapse under the natural arch is obtained from the formula:

-x-h'-*(«-*)]-

b

<

l

H

where a, h = dimensions of the rectangular opening, φρ = angle of internal friction of the rock, H = distance of the roof below the ground surface, fp = coefficient derived from the cube strength of the rock l-j7wr L corrected for the estimated effects of Assuring and weathering. Values of fp and
316 TABLE 20-1 Height of rock load over an opening (according to Terzaghi) Category of rock

Rock load height (v)

Behaviour of rock

A - - massive rock

0—0.256

possibility of larger bursts and falling of smaller fragments, no lateral pressure

B - - bedded rock horizontal beds vertical beds inclined beds

0—0.500 0—0.256 0.25—0.506

falling of stones, no pressure

C - - irregularly jointed rock

0.25—0.35 (6 4- h)

falling of stones, lateral pressure small or absent

0.35—1.10(6 + h)

unstable roof, lateral pressure small to considerable

E - - cohesive soil of medium depth

1.10—2.10(6 + h)

unstable roof, large lateral pressure

F - - deep cohesive soil

2.10—4.50(6 + 6)

unstable roof, large lateral pressure

D - — densely jointed rock, crumbly rock non-cohesive soil

lateral

Caverns are usually designed with vaulted ceilings, so that the major part of the loosened rock below the natural arch is removed in the excavation. In caverns which have already been excavated, the zone of the natural arch can be ascertained approximately by any direct measuring method (for example geophysically). If the theoretical height of the natural arch is known, the optimum length and spacing of the bolts can be roughly assessed. The purpose of the anchorage is either to suspend the loosened rock of the cavern roof from the loadbearing zone of the natural arch, or to strengthen the loosened rock so that it becomes self-supporting, further loosening of the rock thus being prevented. The former type of anchorage which draws the excavated face in towards the zone of the natural arch, can be used if the arch zone is not too far in from the cavern face, and comprises strong and little-damaged rock. The maximum bolt length required is usually that of the bolts placed in the centre of the roof, and is equal to the sum of the distance between the lower edge of the natural rock arch and the ceiling, and the fixing length of the bolt in this zone:

317 I = υ + hu

or

v + /„,

where /„ is the fixing length of the bolt in cement. The spacing of the bolts, / r , is determined, as in the preceding Section, from the bolt's load-bearing capacity and the weight of the suspended rock at the point of the longest bolt:

/,- /SIS,

V where Fs = xt = y= v =

ν.γ area of cross-section of the bolt, permissible tensile stress of the bolt material, volume weight of the rock, loading height.

For practical purposes and for the sake of economy lr is usually taken as being at least 1 m, and the diameter of the bolts is adjusted accordingly. In the second case, which applies in the presence of weaker or damaged rocks and where the natural arch is formed farther in from the excavation face, the rock in the loosened zone is reinforced and prestressed by bolt anchorage to form a load-bearing arch, d [166] (Fig. 20-7).

Fig. 20-7. Artificial arch formed by locking the loosened rock above the opening with a system of prestressed bolts / / A — width of arch formed where — = 3, B — width of arch formed where — = 2 lr

lr

If the apex angle of the pressure cones issuing from both ends of a bolt of effective length / is 2a, and the bolt is prestressed with a tensile force P, the compressed zone, d is subjected to a radial stress from each bolt [69] r given by the expression: 8P Gp

~

nl2Ag2a'

This radial stress induces a peripheral stress in the rock mass; the latter

318

stress acts in a direction perpendicular to the axis of the bolt, and substantially increases the strength of the rock in the compressed zone [113]. It has been found from triaxial tests that a rock, the strength of which was very low under uniaxial compression, acquired a strength of 2 to 8 MPa under a lateral compression load of 0.2 MPa. An arch formed from rock which when prestressed acquires a strength of 5 MPa in a zone 100 cm wide, can take a peripheral load of up to 50 kN/cm, and is thus equivalent to a concrete arch 20 cm thick [69]. The width of the compressed zone, d, with effective bolt length /, bolt spacing / r , and pressure cone apex angle 2a = 90°, may be approximated as d= I - lr. In this way, a continuous compressed zone of rock is created of sufficient thickness to transfer the dead weight of the loosened section of the overburden. The length of the bolts is often determined empirically, as described in Section 20.2.2. 20.1.3

Effect of natural planes of discontinuity

The natural planes of discontinuity which always occur in strong rock, although with different densities and different degrees of regularity, may considerably affect the stability of a cavern and must therefore be taken into consideration in the design of anchorages for the face of the cavern. If one system of discontinuity planes, such as stratification, is markedly developed in the rock, the orientation and position of the bolts have to be adjusted to the natural conditions (Fig. 20-8). The mounted bolts should traverse these planes of discontinuity either perpendicularly or at an angle of at least 90° — φ, in order to increase sufficiently the resistance to mutual displacement along the planes, and enable the rock to transfer compressive forces both along and across the planes, in the manner of an independent arch. The magnitude of this additional resistance within the area secured by one bolt must be such that the total resistance to movement between planes of discontinuity must be greater than the forces (usually the dead weight

Fig. 20-8. Placement of bolts for various directions of the planes of discontinuity A — horizontal, B — vertical, C — slanting

319 of the rock mass) which tend to bring about this movement. If the angle of inclination of the main system of planes is a (see Fig. 20-8), then the maximum tangential force, Γ, which must be secured is given by T = G . sin a = = G cos a . tg φ + Rk = y . V . /r2 . cos a . tg φ + Rk, and therefore, Rk = y . Γ . /r2(sin a — cos a . tg φ), where y /' lr φ

= = = =

volume weight of rock, effective length of bolt (/' = v or /' = /?s), spacing of bolts, angle of friction along the plane discontinuity.

The additional resistance, Rk, to movement along the discontinuity is usually created by a tension, Pk, in the bolt. If the bolt forms an angle ψ with the horizontal (see Fig. 20-8c), then in a vertical section perpendicular to the plane of discontinuity, Rk = Pk . cos (a — ψ) . tg φ + Pk . sin (a — ψ). Substituting for Rk above, _ k

y . /'. / 2 (sin a — cos a . tg φ) sin (a 4- xj/) + cos (a + i/0 . tg . φ '

Furthermore, where Fs = area of cross-section of the bolt, xt = permissible tensile stress of steel. The force Pk will be most efficient (creating maximum Rk) when the angle of inclination of the bolts, φ, is α + φ — 90°. For horizontal beds and vertical bolts, Pk = —G.tg(p. A detailed analysis of the equilibrium conditions in a block of rock near the walls or ceiling surface of an underground excavation was worked out by Lang [113] on the basis of a two-dimensional stress system. He considered a block in the ceiling or wall, intersected by one or two joints, and subjected to (apart from the effect of its dead weight) the resultant of external forces and the compression effects of prestressed bolts in various directions. The conditions of equilibrium for the block of rock in the roof and wall of the excavation are shown in Fig. 20-9. At the joint surfaces, it is assumed that the only resistance to movement is that induced by friction. According to Heuze [77], there is still another resistance to displacement along the joints in strong rocks, namely that occurring as a result of the unevennesses of the joint surface; this effect would have to be overcome by the dilatancy of the rock (lifting, increase of volume) before any sliding could take place along the joint. The difference between the shear resistance at the

< tg (α + ψ) B : tg (χ>

tg (a + φ)

Fig. 20-9. Conditions of stability of a block of rock at the face of an opening. A — in the ceiling, B — in the wall [113]. The block is divided by two planes of discontinuity (joints) and is subjected to the dead weight of the rock, G, the resultant of the forces exerted by the surrounding rock mass, Ry and the tension of the bolts, P

displacement mm

Fig. 20-10. Comparison of peak shear strength for a non-dilatant joint (/), with that for a dilatant joint (//), with the same peak angle of friction [77]

joint which does induce dilatancy, and that which does not induce dilatancy on shearing, is shown in Fig. 20-10. Prestressed bolts are an effective restraint on rock dilatancy. If non-prestressed bolts, cemented in the boreholes along their entire length, are used for roof stabilization, they resist movement along the discontinuity by virtue of their shear strength. In this case only the permissible shear stress, κβ, of the bolt steel is considered: k

cos (α + φ) ' and the required cross-section of the non-prestressed bolt (to provide the necessary shear strength) is given by: _ y . /' . /,(sin a — cos a . tg φ) cos (a + φ) If the angle of friction along the planes of discontinuity is small, nonprestressed bolts embedded in grout along the entire borehole length offer greater resistance. Such bolts offer resistance only when the rock strength is exceeded. 20.2 A N C H O R A G E D E S I G N FOR U N D E R G R O U N D EXCAVATIONS

The stability analysis and anchorage design for an underground excavation must take into account the geometry of the excavation, the geological structure in the wider surroundings of the excavation, the physical and mechanical characteristics of the rock mass including its initial state of stress,

321 and the excavation method. These starting conditions may vary from the simple to the very complex, and from situations which are well understood at the outset to those that can only be more accurately understood as the excavation proceeds. The anchorage design is decided by these starting conditions, which are known either empirically or from the results of analysis; thus the design may be complemented, if necessary, by the results of observations and measurements carried out in the course of excavation. Even where rule of thumb is followed, preference must always be given to a general pattern of bolting, rather than to the installation of bolts only where the engineer or inspector considers they might be needed [61]. No jointing pattern in the rock is completely regular, and therefore the positioning of bolts only according to superficial surface conditions could well have disastrous consequences. Pattern bolting has many advantages from the construction point of view. The number of bolts in each row and the distance between each row are arranged so that it is possible to install one or more rows after each round (cut) is fired, and the installation crews can then work systematically and quickly. With such systematic working, each bolt is assured greater attention both during its installation and its subsequent checking. These advantages far outweigh the cost of the few extra bolts that may be used in this way. The design must in every case be based on up-to-date geological information concerning the location of joints and other major features of the geological structure, and there must be a careful appraisal of the information obtained from rock behaviour measurements in areas that are already opened up. A very important part of this appraisal is the observation of the effect of deformation on the loading of the excavation and the change of this loading with time. Immediately after excavation of the cavity an elastic displacement of the rock into the free space takes place together with a marked drop in radial pressure. The time of this occurrence is also the most suitable time for setting up the support and reinforcement of the rock, because a relatively weak reinforcement then suffices for its stabilization [175]. Delay in placing the reinforcement leads to a gradual loosening of the rock in the surroundings of the excavation and a further increase in pressure. An inadequate reinforcement then has to be strengthened, and above less deep excavations, the ground surface may sink [35]. 20.2.1

Analytical procedure

According to Gerhart [61] the analytical methods used for assessing the stability of rock structures are direct developments from structural analysis and applied mechanics. Their complexity ranges from the simple case of

322

a block sliding on a plane surface of known frictional resistance, to highly complex finite element solutions that include the effects of slippage along discontinuities and the fracture of reck blocks. The usefulness of any analytical solution is determined not by its arithmetical accuracy, but rather by the accuracy with which it represents the parameters of rock mentioned in the introduction to Section 20.2. Analytical methods are basically divided into two types, those including elastic analysis and limit analysis, which perhaps is more valuable in the design of rock reinforcement. Elastic analysis techniques currently in use are: a. methods of calculating stress concentrations around openings, b. finite element method, c. laboratory model methods using optically sensitive materials (photoelasticimetry). Several authors [97, 149, 143, 200] have already presented solutions for calculating the stress conditions near single and multiple openings in stressed elastic media. Goodman and Heuze have given useful supplementary design tools for finite element analyses. Limit analyses for rock structures have been presented by many authors, and simple examples of this have been introduced in Section 20.1. At present, the finite element method is most often used for the stability analysis of underground excavations. It takes into account, with minimum difficulty, many of the factors which affect the stability. The calculations, obviously, can only be done with the aid of a computer. Analysis by the finite element method is based on the assumption that the surroundings of an underground excavation can be considered as a great number of small geometric elements with three or four apices (Fig. 20-11), increasing in size with distance from the opening as their effect on the stress state diminishes. The calculation considers a unit displacement of one apex of the element and the force which can be said to have induced the deformation of the element is sought. This force must be equal to the resultant of all forces actually acting on the rock element. Physical and mechanical characteristics of the rock mass, found as a result of investigation, are substituted into the deformation equations set up for one element. Similar equations are obtained for the other elements of the net. The entire program comprises a system of several hundred to a thousand equations. The computation gives the magnitude of the stress at different points in the excavation surroundings. By studying these points, any zone in which the strength of the rock might be exceeded can be identified. The effect of prestressed anchors is then introduced into the equations as an external loading of the rock, and the procedure is repeated, giving an indication of how this effect influences the state of stress in the surroundings of the excavation, and to what extent the dangerous stresses are brought below the admissible limit.

323

Fig. 20-11. Distribution of finite elements in the surroundings of a rock opening

kx>

A**TZ

As an example, the stability solution obtained by this method for the underground cavern of the Machu Picchu power station (Peru), is shown in Fig. 20-12. The nets of elements are laid out in cross-sections so that the effect of two main systems of planes of discontinuity in the granodiorite, and the effect of the stabilizing anchors in the roof and walls of the cavern are shown up by the analysis. 20.2.2

Empirical procedure

When neither the geological conditions nor the scale of the operation justify the use of exacting analytical procedures, anchorages for the stabilization of underground excavations can be designed according to empirical rules based on experience. Many kilometres of tunnels all over the world have been successfully built in this way. Also the well known New Austrian tunnel driving method, which is described in detail in the following Section, is based on a rule-of-thumb anchorage design according to qualitative evaluation of the natural rock conditions. In Europe, the most widely used empirical rule for the design of bolt anchorage in tunnels was laid down by Rabcewicz in the fifties [172]. He recommended that the effective length of the bolt should be equal to, or greater than, one third of the excavation width, and that the spacing of the bolts should not exceed a half of the effective bolt length. The prestressing of the bolt should equal approximately the weight of the secured rock.

324

A

0 I

5 '

10 15 '

»

20

1 m

Fig. 20-12. Finite element mesh for Machu Picchu underground power station in Peru [42]

In the USA, a detailed set of empirical rules governing the length, spacing and prestressing of bolts was laid down by the Corps of Engineers [61]. They recommend the following parameters: Length (minimum) a) Two times the bolt spacing. b) Three times the width of unstable rock blocks. c) For roof bolts spans less then 6 m —one half of span, spans from 6 to 18 m —interpolated within the range 3 to 4.5 m bolt length, spans 18 to 30 m — one quarter of span

325 d) For wall bolts height less than 18 m —length as determined in c) above, height more than 18 m —one fifth of height. Spacing (maximum) a) one half of the bolt length, b) one-and-a-half times the width of unstable rock blocks, c) 1.8 m (a spacing of more than 1.8 m makes the attachment of a surface net such as chain-link fabric difficult). Prestressing (minimum average confinement pressure at yield point of bolts) 1. For roof bolts a) pressure equal to that of a rock load of vertical thickness 0.2 times the opening width, b) pressure of 42 kPa. 2. For wall bolts a) pressure equal to that of a rock load of vertical thickness 0.1 times the opening height, b) pressure of 42 kPa. 3. At intersections of underground passages Twice the confinement pressure as indicated above. This reinforcement should be installed in the first passage (opening) excavated prior to forming the intersection. Stress concentrations are generally higher at intersections, and rock blocks are free to move towards openings. 20.2.3

New Austrian tunnel driving method

This method has come into use all over the world during the last two decades, although it is based mostly on empirically derived knowledge. It is not new as regards the driving procedure, the main innovation being the method of securing the excavation by the stabilizing effect of anchors. The rock in the surroundings of the opening, damaged and loosened by the excavation work, is strengthened by a regular system of steel bolts to form a self-bearing, but yielding, roof arch. The bolt system is complemented at the rock surface by a layer of gunite of varying thickness, reinforced by wire mesh or steel ribs, if necessary. This reinforcement can be adapted for either temporary or permanent stabilization of underground excavations of a variety of cross-sections; it can be used in full face tunnel sections or in parts, while explosives, tunnelling machines or shields are being used nearby. The extent of the strengthened zone around the excavation can be varied according to the quality of the rock and the outline of the opening. This zone can easily be strengthened with further anchors or layers of gunite, if such seems necessary on the basis of deformations of the rock and rock reinforcement

326

registered by instruments set up in the course of excavation. The reinforcement is quickly installed with a high degree of mechanization, made possible by the fact that the opening remains free all the time. The full opening usually has a circular or horseshoe shape. The design of the anchorage and its complementary strengthening is usually carried out according to a standard scheme, there being groups of such schemes corresponding to particular qualities of the rock or soil. In Europe, the classification of standard schemes compiled by the Austrian experts Rabcewicz, Lauffer and Pacher [114, 161] is well known. There are six classes with corresponding construction sequences and reinforcement (Fig. 20-13). 1st class. Massive, unjointed, or slightly jointed dry rocks the compressive strength of which is sufficient to withstand the tangential stress in the excavation line. The complete excavation is permanently stable without reinforcement, or with minimal local strengthening of individual rock blocks, or places susceptible to bursting; for the latter purpose, short bolts are fixed individually or in groups. /.

//.

///·

Fig. 20-13. Six classes of tunnel excavation scheme, with corresponding support construction sequences (New Austrian tunnelling method, according to Pacher) [161]

2nd class. Rocks penetrated by a network of planes of discontinuity (bedding joints, vertical joints). Water seepage is not great, and the rock strength in the excavation line is not exceeded. The full tunnel section is excavated, and permanent stability is secured by a regular system of anchors

327

in the roof, together with wire mesh. The walls and floor are locally strengthened with anchors as required. 3rd class. Rocks densely or very densely dissected by planes of discontinuity in different directions (stratification, foliation, and/or jointing). Crushed zones and clayey infillings are present and there is visible water seepage. The rock strength in the excavation line is exceeded and the rock must be systematically strengthened to form a load-bearing roof arch around the excavation. The zone of loosened rock above the ceiling is threatened by collapse first of all. Excavation proceeds in two stages, first the roof section and then the floor section, with immediate securing with gunite initially, then with anchors or steel girders, and subsequently with gunite again. Prestressed bar anchors can be used for this category of rock, but after prestressing they must be fixed in the rock with cement over the entire borehole length. The more the quality of the rock mass has deteriorated, the more do prestressed anchors have to be replaced by non-prestressed anchors fixed in the rock along their entire length. 4th class. Badly broken to technically crushed rocks, regions of rock failure and cohesive soils of stiff consistency. Plastically deforming rock or soil intrudes spontaneously into the excavation from the roof and walls and the floor rises. There is marked water influx. Excavation proceeds in several successive stages, always with immediate securing with anchors, steel girders and gunite. The strengthened rock zone must be completed with an adjoining concrete vault at the bottom. The class 4 standard anchored reinforcement scheme used for the Taurus motorway tunnel in Austria, is shown in Fig. 20-14 [69]. 5th class. Crushed, mylonitized rocks, cohesive soils uncompacted, much squeezing (pressure-exerting). Plastically deforming material intrudes into the excavation from all sides. There is a considerable water influx. Securing is achieved in the same way as in the previous class, except that longer anchors are used. Excavation and anchoring sequences are shown in Fig. 20-15. 6th class. Loose soils, detritus and crushed rocks at great depth, and generally the most difficult conditions for excavation. Excavation progresses in short stages analogous to the sequence for class 5, or a shield is used; however the length and density of anchors is greater, and the spacing of steel girders smaller. Even the front face of the excavation must be secured with gunite or anchors. The bottom of the excavation is also anchored to create a conjoined load-bearing arch, which must, however, show a sufficient degree of yielding. The rock pressure is significantly reduced if a small yielding of the reinforcement is possible. If, however, excessive deformations of the rock are registered in the surroundings of the opening (Fig. 20-16), the reinforcement is strengthened with further anchors.

328 Fig. 20-14. Standard support system for the 4th class of rock (according to the New Austrian tunnelling method), as applied in the Taurus motorway tunnel in the Alps. Anchors installed in the sections with strong lateral pressures are shown by dashed lines [69]

cross-section 0.15

A-A , 0-15

concrete öottom vault

0.75ml

\ steel laggings > wire mesh A 65 \ shoterete 26 MPo steel niös TM 36/58 isolation inside concrete lining

Fig. 20-15. Excavation and anchoring sequences for the 5th class of rock, according to the Austrian classification [69]

329 Fig. 20-16. Installation of measuring equipment in tunnel excavation, according to. Müller [143] A — 1 — prestressed anchors with load meters, 2 — multiple position extensionmeter, 3 — section between studs fixed in the rock face for the measurement of the convergency of the excavation B — measurements of convergency in the Taurus tunnel by workers of Interf els, Salzburg;

330

The New Austrian tunnel driving method has proved its worth not only in strong rocks, but particularly also in squeezing and loose ground where astonishingly good results have been obtained. For instance, during the construction of the Massenberg tunnel (Austria) [174] in slope detritus and weathered shales, caving-in occurred even with strong concrete reinforcement 80 cm thick; the rock was eventually stabilized by the use of anchorage. Another example of the successful application of this method in very adverse conditions is that of the construction of the underground railway in Frankfurt (GFR) [69]. The tunnel of diameter 6.35 m was driven through cohesive soils with a maximum compressive strength of 0.3 MPa, angle of friction 20°, and cohesiveness 10 — 65 kPa. The overburden was only a few meters thick and the tunnel passed the foundations of buildings at a distance of only 6.20 m. The geological data necessary for the design of the anchorage system were obtained from an exploratory gallery, driven in advance. The faces of the full excavation were anchored with particular care to minimize settlement of the ground surface. The New Austrian tunnel driving method is very adaptable to new conditions of the rock mass met with in the course of excavation, and the reinforcement can be strengthened almost arbitrarily, if necessary. For example, when sections of the Taurus tunnel were driven at a depth of 800 to 1,000 m into highly compressed phyllites originally placed in class 4, the reinforcement turned out to be inadequate. Large deformations rapidly developed during excavation of the roof section as a result of large lateral pressures. To deal with this, further non-prestressed grouted anchors 6 m long (and later 9 m long) were installed, and ultimately two rows of 13 m-long rope anchors, prestressed to 600 kN had to be added. Only then was the section stabilized. Most of the communication tunnels all over the world are driven by this method nowadays.

20.3 EXAMPLES OF T H E A N C H O R I N G OF U N D E R G R O U N D EXCAVATIONS

20.3.1

Anchoring of the roof of an excavation

The first part of an underground excavation to be secured and stabilized is its upper part, that is, the roof, because collapse of the rock under the force of gravity occurs most easily, and therefore most frequently in the roof. This procedure was adopted in the classical tunnelling methods, and is followed in present-day excavations of full sections. Rock anchoring is a rapid and very efficient means of roof stabilization.

331

The basic anchorage parameters for an underground opening (gallery, tunnel or cavern) were considered in the preceding Sections. The distance in the longitudinal direction between individually anchored cross-sections is often equivalent to the bolt spacing within a cross-section (s = lr); this distance is greater if steel bands are used in the longitudinal direction. Deciding on the reach of each cut during driving is also important, and depends on the rock type and the available equipment. The distance of cut, z, is often set about the same as the bolt spacing in the opening cross-section. This distance of cut is thus related to the standing capacity of the rock with time and the excavation width, as verified in previous work or from geotechnical surveying or the information given in Table 20-11. The latter is based on Austrian experience of present-day tunnel cross-sections up to a frontal area of 100 m 2 [114]. This classification was simplified by Bieniawski [41]. TABLE 20-11 Approximate period of stability, and spacing of supports in tunnels in different types of rock (according to H. Lauffer) Type of rock

Period of stability

Maximum spacing between supports

A — strong rock (compact, igneous rock, massive thickbedded sediments, massive gneiss) B — jointed rock (jointed igneous rock, thin-bedded sediments, metamorphic rock with marked foliation) C — densely jointed rock (densely jointed igneous rock, shales and weaker metamorphic rocks) D — crumbly rock (soft rocks, clayey shales, disturbed and partly weathered hard rocks) E — very crumbly, disturbed rock (weaker soft rocks, much disturbed and weathered hard rocks) F — pressure-exerting rock (weathered and disturbed clayey shales, cohesive soils with solid to hard consistency, sand and gravel with high moisture content) G — high-pressure-exerting rock (cohesive soils with soft to stiff consistency, saturated sand and gravel, fills, organic

20 years

4 m

6 months

4m

1 week

3m

5 hours

1.5 m

20 minutes

0.8 m

2 minutes

0.4 m

10 seconds

0.15 m

soils)

When a new cut is made, a rock arch or beam is assumed to come into effect between the last line of bolts and the excavation face. Over the newly broken roof section, between the points of support, a normally parabolic

332

zone of loosened rock develops, with a height not exceeding —. A new line of bolts is placed at the stated distance lr to secure the roof. By tightening the nut on each bolt the washer is pressed against the rock surface with a force of at least 30 — 40 kN, and this prestressing strengthens, stabilizes and prevents further loosening of the reck in the close vicinity of the anchoring point. When the rock is densely fractured, there is the danger that rock fragments may fall from the roof in between anchoring points, or in massive rocks, bursts may occur as a result of concentrations of stress in the excavation face. In such cases protective wire nets are laid along the roof surface immediately after its first rough dressing, which can be carried out from the preceding, already secured section. The fixing washers of the anchors press the wire net against the rock surface (Fig. 20-17). Steel bands, connecting several bolts in a row, and used in place of the individual washers, provide a high degree of stability. A layer of gunite or concrete sprayed on after the anchors have been prestressed provides a lasting protection of the surface (Fig. 20-18 and 20-19).

Fig. 20-17. Wire mesh on the surface of an excavation secured with anchors (photo Goldenberg)

333

Fig. 20-18. Application of gunite to the face of an excavation (a) and a machine for shotcreting, guniting and pneumatic conveying Meyco GM 060 of Intradym AG, Switzerland (b)

S*ÄC

Ψ ;Xf''V> ? ;', '■'&;„']

a)

*>'-'/: ^^mm^^mM *j*jT ^

\

*>) The whole cutting cycle, including drilling, placing the explosives, blasting, dressing the roof, removing the spoil, and anchoring the roof, should be organized in one working day. When suitable machines are available, particularly multi-purpose wagon-drills and loaders, the cutting may advance across the full cross-section (Fig. 20-20), even though this may be very large (over 100 m 2 ). Thus, for example, when a railway tunnel in Norway (area of cross-section, 70 m 2 ) was driven into strong granite, the length of cut9

\

\

\

■

IM

ί j

T

bolting floor

f^f

^ f

^Är

Fig. 20-20. Excavation sequence in the Gotthard highway tunnel in Switzerland [66]

*4 Fig. 20-19. Excavation of a gallery (diameter 5.8 m) for the Suassaz hydroelectric power plant (France). The excavation is secured with a regular array of bolts covered with a layer of gunite [212]

335

and therefore the daily advance, was 3.60 m. Eighty boreholes were made in the advancing face, the peripheral boreholes carrying limited amounts of charge so as to obtain a smooth blast face. A Jumbo wagon-drill with four drilling booms of Atlas Copco advanced 50 cm/minute using a 1 7/8" (48 mm) bit [74]. The entire roof anchoring operation across the full width of the excavation is now highly mechanized. Several firms produce special Jumbo wagondrills which are remotely controlled from the previously secured section. Such a machine (Fig. 20-21) automatically drills the anchoring boreholes whether vertical or inclined, places the mechanical or grouted bolts and then prestresses them by tightening the nut to the required tension. Only 3 minutes are required for the fixing of one bolt into the roof. The smallest existing self-propelled single-boom hydraulic jumbo Secoma ATH 12-1F can drill and bolt in galleries as small as 2 x 2 m and as large as 3,9 x 4,2 m. For higher openings, other types of wagons are manufactured with telescopic platforms (Fig. 20-22) from which the roof is dressed, anchored, and covered with wire net or mesh after the blasting. 20.3.2

Anchoring of communications tunnels

As mentioned earlier, the majority of tunnels for highways, railway lines, and urban underground ways are now driven and secured by anchored reinforcement according to the principles of the New Austrian tunnel driving method. Some examples were described in Section 20.2.3, and another example is shown in Fig. 20-23. In the Katschberg tunnel, anchors were installed to one side of the tunnel to strengthen the excavation where it is cut into shales (Fig. 20-24). The remaining part of the tunnel section, cut into gneiss, was stable without reinforcement. The anchors were of the mechanical GD type, with a diameter of 16 mm and a synthetic base; the reinforcement was complemented by wire mesh. Communication tunnels very often are driven by non-destructive methods in which partial- or full-face cutting machines are employed. In this progressive method, it was impossible to secure the rock face by anchors immediately following the excavation, but only at the rear of the cutting machine; this caused a loss of time, which could impair the stability of the excavation. At present, for example the cutting machines a boom cutter or a full-facer are being equipped by the Austrian Böhler Co., with a drilling attachment (Fig. 20-25), allowing to anchor the excavation directly at the rear of the cutting head. This drilling attachment is useful also in blasting operations, when hard rock is encountered at the front of the excavation. Tunnels of larger cross-section, and tunnels built in difficult geological conditions, are driven in parts divided by working faces. Thus, for example,

336

a)

Fig. 20-21. Bolting jumbos of Tamrock BH 20-8 (a) and Secoma CTH 15-1B (b) for fully remote-controlled rock bolting

the Gothard motorway tunnel in Switzerland had an excavation width of about 11 m and was mostly driven in full cross-section (see Fig. 20-20) with simple ancorage of classes 2 and 3. However a short section 320 m long in paragneiss turned out to be very difficult; it was necessary, first of all, to

337

Fig. 20-21.

drive and reinforce narrow galleries around the perimeter of the full cross section (Fig. 20-26). Because a reduction of the cross-section width by up to 150 cm took place under the very high lateral pressures, more than 800 anchors from 6 to 9 m long, prestressed to 580 kN, had to be installed in this section. The final concrete lining of the tunnel could only be carried out after gradual disappearance of the rock pressure. The stabilizing effect of anchors was exploited in a remarkable way when the wide stations of the Washington underground railway were excavated at a short distance below the ground surface [35]. The stations, excavated to a width of more than 20 m, are mostly less than 30 m below the ground surface, and the rock cover over the crown is often less than 10 m. The schistose gneiss in which many of the stations are cut is unweathered, and contains four or five sets of joints which are planar, continuous, and often smooth-faced. The joint spacing commonly ranges from 1 to 2 m. The shear

338 Fig. 20-22. Wagon-drill with telescopic platform, in service, A — anchoring the face of the cavern for the Churchill Falls hydroelectric power plant in Labrador (photo Williams)

«ill.« 1

\Mmmm

« i f f If»?·»

'"■/'"

wmwM i ί

B — laying of wire mesh on the face of an excavation already covered with a layer of gunite in the Monte Piazzo Tunnel in Itally (photo Titanite)

zones which strike parallel to the foliation are typically 0.5 to 2.0 m wide and consist of layers of fractured rock with smooth interfaces. The quality of the rock apart from the shear zones is usually high (RQD = 70 % or more). At first the inter-station tunnels of 6 m diameter were excavated by

339 cross - sectional

^ 777^?Zyr^77r^7/>

luMjitudinel section

Fig. 20-24. Strengthening by means of short bolts a part of the roof formed by shales in the Katschberg tunnel (Austria) [69]

io

bolts 3.5m < ) 2 layers of J shoicreie with mesh Fig. 20-23. Support for the Monte Piazzo tunnel, according to class 3 of the New inside concrete 1

11

^

12

1.1

Austrian tunnelling method [212]

T-*

^R;

o)

Fig. 20-25. Two hydraulic booms HB 450 of Böhler with drilling and bolting equipment mounted on a boom cutter a) — schematic drawing, b) see page 340

tunnel-boring machines cutting through the station areas as well. The openings for the stations were excavated and secured in a later phase. Fig. 20-27 illustrates the construction techniques which were developed for some of the stations. Inclined bolt anchors were installed from the previously excavated line tunnels to provide protection for the future station aich as the upper heading was excavated. This pre-support system permitted subsequent excavation stages to be of greater width than would otherwise have been possible. During the initial stages the sidewalls were also secured by casting

340

II,

b)

Fig. 20-25. 6) — view of a cutting machine with two booms of Böhler in idling position

Fig. 20-26. Excavation of the Gotthard tunnel in parts (galleries) in a difficult section in paragneiss a) — excavation scheme,

341

b) Fig. 20-26. b) — view of excavation [66]

the wall plate and the lower portion of the concrete arch. The final structural lining, consisting of steel ribs and gunite, was installed as the heading of the fullwidth excavation was advanced (Fig. 20-28). Detailed measurements of rock deformation made with extensionmeter demonstrated the high efficiency

342

typical joint orientations

recessed rock bolt

Fig. 20-27. Excavation sequence for the stations of the Washington Underground [35]. Bolt spacing: 1.5 m along the tunnel

foliation shears

1a,b running tunnels ; 2 pilot tunnel

initial

Ί

^rib W n*61

shoterete layer

sho terete

Fig. 20-28. Typical dimensions of the surface lining in the Washington Underground stations

of the anchorage, even under the extremely adverse conditions prevailing in the Dupont Circle Station (Fig. 20-29), where the line tunnels had not been excavated in advance. 20.3.3

Anchoring of small openings, rock pillars, galleries and shafts

The anchoring of smaller openings and rock formations in the course of constructing underground systems is usually based on empirical procedures. Short bar anchors or bolts, without prestressing and with their entire length embedded in grout or resin, are used to reinforce the superficial zone of the rock, for which purpose they are very effective even in swelling rock. The securing of flat ceilings in horizontally bedded rocks (Fig. 20-30) is a typical example of the use of such anchorage. The anchored excavation roof in the White Pine copper mine (Michigan) is shown schematically in Fig. 20-31 [69]. The width of excavation achieved by securing with short bolts in the overlying beds of sandstone is remarkable. The application of anchoring to the roofs and faces of inclined openings and haulage shafts (Fig. 20-32) is particularly appropriate, since the construction of any other type of support in these situations is very difficult. In stabilizing the greatly stressed surfaces of hauling drifts and shoot holes*

343 multiple position

Connecticut A ve underpass

Fig. 20-29. Instrumentation (A), excavation and support construction sequence (B), in the Dupont Circle Station

cluster ofdouble V strain gauges attached to position ettensometer steel sets

rock bolts f 28 mm

Jo 0f\^

3cr>sho terete 3b-±\ s|

/

/

bolts^z f35mm

concrete

^b

/ 6b 0

1 i

i i i

5

1 i

10m

^

I —

i i ij

bolts and ribs spaced 1. 52 m longitudinally ribs WMx61,shotcrere 15to 60cm thick

the best results are obtained with non-prestressed bars without externally projecting heads which might obstruct the passage of spoil etc. Cements based on synthetic resins are particularly suitable in these situations, as they remain sufficiently elastic after hardening to absorb and withstand the severe shocks produced by explosives and movement of disintegrating rock [130]. The anchoring of the concrete linings of pressure galleries and penstock shafts fulfils a different function, namely that of combining the strength of the lining with the strength of the surrounding rock against the internal pressure of water. A reinforced and anchored gunite lining (see Fig. 20-19) can be used successfully in place of more costly steel armouring. Very effective strengthening by means of prestressed bolts and anchors can be achieved in rock pillars and in corners, where the anchors assist in distributing high concentrations of stress throughout the rock of these forma-

344

Fig. 20-30. Anchored flat roof of a gallery in underground mine (photo Titan, Australia)

tions. A good example of an anchored rock pillar is given by Polish authors [201] (Fig. 20-33). In cases of swelling rocks, non-prestressed bolts have proved to be very efficient [172]. Thus, in the experimental section of a gallery 2.40 m wide in which the floor was pushed upward by 1.40 m, the floor level was restored to 10 cm above the original level after four bolts 1.50 m long were fixed in successive cross-sections 0.90 m apart. Another example of the stabiliza-

345

οσΤ

Fig. 20-31. Strengthening of roof bed with bolts in a wide opening of the White Pine mine (Michigan) [69]

:.?ϋΚ1ΐΐηΜ 9.0 m

-A

section A -A

3m

Fig. 20-32. Reinforcement of a boxhole with non-prestressed bolts fixed with resin

tion of a swelling clayey shale forming the floor of a coal mine gallery is shown in Fig. 20-34 [69]. In these situations the bolts must be fitted at the earliest opportunity after excavation, the anchoring must be of sufficient depth (1/3 to 1/2 of the gallery width), and the grout must be of a rapid hardening type. 20.3.4

Stabilization of large underground caverns

Large underground caverns are constructed for various purposes, often to contain the equipment for hydro-electric power plants and also as special storage for materials such as oil products and foodstuffs. Caverns may be excavated for the installation of military equipment, for sport halls, cinemas and churches, etc. This widespread underground construction method, even in places where the rock does not have the most favourable characteristics,

346

i

wmi §§§■■

w

mm

9* '-•.Λ

Ji —v&e^ß'i'w 4-

.

b>r

«s

'

Γ^ v

-·*£/*'

t * * , l£&h,~*&.

,*M%ippi

Fig. 20-33. Anchoring of the upper part of a rock pillar 6 m high iii the Olkusz mine (Poland) [201]

has been made possible by the technique of rock anchoring. As long as only supporting types of reinforcement were available, the excavation of large caverns in weaker types of rock was difficult and very- costly. Nowadays the excavation of large spaces proceeds quickly and is economical. Under-

ground

plan

^J

fc

S* -o

o

^

1

-o

Ά

^

o-

fcs

o

o

-o———o

-i

o-

4.2/77 o~

"if 5k

®

cross - sect/on

Fig. 20-34. Successful anchorage of a gallery floor consisting of swelling clayey shales in the Hugo coal mine in the Ruhr (GFR) [69]

ground caverns are much cheaper than surface constructions of the same usable volume. Among the first large, anchored underground caverns in the world was the machine hall of the power plant at Lipno in Czechoslovakia (Fig. 20-35), the maximum width of the arch being 32 m [230]. The project involved an investigation of the structural conditions of the granite mass, carried out by geophysical surveying in pilot tunnels. As a result, faces of up to 30 m in height in the main cavern were reliably secured with anchors. Steel bolts, 36 mm in diameter and from 4 to 9 m long, were installed in boreholes filled with grout, the grout having been passed to the blind end of the borehole by means of a glass tube. When the grout had hardened, the bolt's were prestressed with a tensile force of between 40 — 50 kN, by tightening the nuts with a flat spanner. The average density of bolt placement was one per 4.2 m 2 (Fig. 20-36). Of the large number of anchor-stabilized underground power plants constructed in the past 25 years, some typical examples are cited in the following. The Lutz underground power station excavated in a flysh series of sandstone, limestone and soft marly shales, is exemplary as a project in which all modern construction methods were used in the securing of the excavation as appropriate for the existing geological conditions (Fig. 20-37) [188b]. During

348

V« ■M%

\>;

w

■*s%

■

■

#

■

;

*

>

>

Fig. 20-35. View of the northern part of the excavation for the main cavern of the underground hydroelectric power plant at Lipno (Czechoslovakia)

349

Fig. 20-36. Typical cross-section of the main cavern at Lipno A — distribution of anchors proposed on the basis of the structural condition of the granite as found in pilot tunnels, B — structural conditions discovered during the full-scale excavation and actual positioning of bar anchors, 1 — lines of intersection where the joint planes meet the section plane (type of joint designated), 2 — anchors, with lengths indicated, 3 — outli™» of the proposed full excavation, 4 — outline of the excavation for pilot tunnels, J> — actual full-scale excavation, 6 — concrete protecting structures

350

the breaking of the roof for the protective parabolic vault, the former was temporarily secured with short bolts mechanically fixed in the rock. The longitudinal walls were stabilized by means of vertical pillars secured with long anchoring reaching deep into the rock, across the beds. The shorter walls, however, were protected from sliding and the collapse of pieces of the rock bedding into the excavated space, by thin reinforced concrete vaults. The Veytaux underground power plant in Switzerland has a cavern 30.5 m wide, 23 to 26.5 m high, and 137.5 m long. It wa& excavated in stages in horizontally bedded and much jointed limestones and marlstones; the perimeter was excavated first, followed by the core of the cavern. The entire face of the excavation was temporarily secured by a regular system of bar anchors, wire mesh, and gunite of 15 cm minimum thickness. The bolts which were 4 m long, were fixed with resin so that it was possible to prestress them (to 160 kN) after several hours. When the excavation had been complet-

Fig. 20-37. Securing of the Lutz cavern (Austria), excavated in a flysch series a) — groundplan (double lines with bedding symbols of strike and dip denote the main positions of soft shales), b) — cross-section of the cavern

ed, the roof and the walls were permanently secured with VSL cable anchors 11 to 18m long, with service loads of 1.35 and 1.15MN. There was an average of one cable anchor per 14 m 2 of roof area (Fig. 20-38). The originally planned concrete arch was abandoned. The underground El Toro power plant in Chile is built in a cavern shown in cross-section in Fig. 20-39. It was excavated in granodiorite with three main systems of joints. The excavation progressed from the top downwards. At each stage of the excavation the rock was immediately secured with long

351

cable anchors (service load, 1,200 kN), together with short bar anchors prestressed to 160 kN, according to the local requirements in regions between long anchors. The spacing of the long anchors was 6 m longitudinally, and 3 to 5 m transversely. When the entire length of the roof section had been

b)

Fig. 20-38. Veytaux underground power plant in Switzerland A — excavation scheme and support construction sequence, B — view inside cavern

352

long anchors · VSL not MR, l°15-17m short anchors: VSL 181 ER,l= 1m

Fig. 20-39. El Toro underground power plant (Chile), showing excavation sequence and anchorage [122]

excavated, a concrete arch 1 m thick was constructed at the crown. This served as an additional strengthening against earthquakes. The Vianden HI power plant in Luxemburg is installed in a shaft 50 m deep and 24,40 en in diameter. The shaft was excavated in a series of clayey shales with Marked foliation inclined at an angle of 57°. A large rock wedge threatening to slide towards the excavation (Fig. 20-40) under its own weight was secured with 102 Dywidag anchors of various lengths passing across the wedge, fixed into the rock mass on either side. Each anchor had a loadcarrying capacity of 1.4 MN and consisted of a bundle of 9 shaped bars 16 mm in diameter with double anticorrosive protection. Mechanical jexpanding bolts (dovetail type, ribbed steel) 9 m long and 32 mm in diameter were used for anchoring the cavern for the Paolo Alfonso IV power station (Brazil), excavated in a complex of hard crystalline rock [58J, The bolts were placed at intervals of 1.5 m to form a grid pattern, and were tensioned to 225 kN with a torque wrench, thus providing a mean compression of 0.1 MPa on the rock surface of the heading (Fig. 20-41). After tensioning, all the bolts were grouted. The first layer of a gunite lining (about 4 cm thick) was then applied. A 10 cm-square mesh of 4.2 mm steel wire was attached to the gunite layer with steel pins, and was tied also to the heads of the bolts. Finally, a second gunite layer was applied to the roof, giving a total lining thickness of 10—15 cm. On the sidewalls of the cavern,

353

a)

lA=20m 15m 10m

Fig. 20-40. Anchorage of rock walls of a circular shaft for the Vianden III power plant (Luxemburg) a) — cross-section and horizontal section, b) see page 354

bolts and dowels (non-prestressed bolts) were placed at regular intervals, and at a later stage, gunite was applied over wire mesh to the unstable regions at the intersections between the cavern and the tunnels. The purpose of the prestressed bolts in the walls was to help stabilize the rock mass by applying a pressure at the rock surface varying progressively from 0.35 MPa in the zones under travelling crane beams, to zero at an elevation of 140.00 m. Bdow this level, only dowels were installed. The crane beams were fixed with 18 m-long anchors of the Freyssinet type, stressed to 1.32 MN.

354

b) Fig. 20-40. b) — view into the shaft

One of the largest caverns in the world (for the underground Waldeck II power station in the German Federal Republic) is 106 m long, 54 m high, and 33.5 m wide. It is situated in a series of clayey shales and greywackes, inclined at 20°, with marked thick-bedded jointing. The compressive strength of these rocks [69, 45, 122] was found to vary from 50 to 80 MPa, and the shear parameters for the bed joints were φ = 20°, c = 0.15 MPa. In view of the dimensions of the cavern, the only feasible method of stabilizing the rock was to construct a self-supporting vault with the aid of prestressed anchors, since an adequate concrete lining would have been far too costly. The stress conditions around the cavity were computed from photoelastic analyses, and the necessary anchoring forces and anchor lengths were determined

355 rockbolts

long tendons 1. (1320kNtensioned)IH

(22 5 kN

tensioned)

elev. -y 151.00 ■ I long tenders ~^j(132QkH tensioned)

rockbolts (22 5 kN tensioned}

unten sionedgrouted dowels(32mm diameter)

unten sioned grouted dowels (32-mm diameter)

(all measurements in metres)

Fig. 20-41. Anchoring scheme and excavation sequence for the cavern of the Paolo Alfonso IV power station (Brazil) [122]

accordingly. The excavation of an oval cross-section of area 1,390 m 2 was carried out in stages from the top downwards, and immediately secured at each stage (Fig. 20-42). The surface of the excavation was provided with a double layer of gunite 20 cm thick, reinforced with wire mesh in each layer. The superficial rock around the entire perimeter was reinforced with Dywidag bar anchors (6 m long in the roof, 4 m long in the walls). The anchors were fixed with resin and were prestressed to 120 kN after 20 minutes. Otherwise, the main anchorage of the cavern consisted of VSL cable anchors of loadbearing capacity 1.7 MN (33 wires of 8 mm diameter, anchor length 23.5 m, spacing 4 m). The walls of the boreholes were tested for impermeability prior to the installation of the cables, and wherever necessary, they were sealed with grout and the holes re-drilled. The anchors and the prefabricated concrete foundation blocks were then placed in position. A few days after the anchors were grout-fixed (fixing length 4.5 m), they were tested to 1.5 times the working force (1.35 MN), and this test was repeated one week later. Ninety anchors of a total of 716 altogether were installed as measuring anchors for long-term observation; the stressed sections were injected with grease instead of grout, and the anchors were fitted with load sensors connected to a central monitoring unit.

356

N

1.7 MN

\

!

/

^ Α ΐ ϊ /

/

/ /

Fig. 20-42. Waldeck 11 underground power plant (GFR). Excavation and support construction sequence d)\ view into the secured cavern^) [45, 69, 122]

The enormous excavation rate of 42,000 m 3 per year in the underground cavern of the Norad Expansion Project near Colorado Springs (USA) [170] was made possible by anchoring. The project required the excavation of three large chambers for the power plant, the cooling tower, and the exhaust valve. The excavation of these amounted to 60 per cent, of the excavation in the entire project (Fig. 20-43). The rock varied from coarse-grained, highly altered and fractured granite, to fine-grained granite. The method of excavation was drill-blasting, using smooth wall and controlled blasting techniques.

357

Fig. 20-43. Situation and geological features of the site of the Norad Expansion Project (USA) [170]

Prestressed, grouted bolts were installed in a regular pattern to provide permanent support. The bolts were of a hollow-core type, with a thrustring mechanical expansion shell (see Section 13.2.4). A pad of quick-setting mortar was built up (35 per cent, failed) in order to seal the orifice round the bolt at the collar, and provide a seating for the plate. The bolt was then tensioned to 110 to 130 kN by means of a straight-pull hydraulic jack, and grouted. BDth of these operations turned out to be excessively time consum ing. 8,664 bolts from 3 to 5.5 m long were used for the primary support (one bolt per 1.3 m3 excavated). Besides these bolts, non-prestressed Perfo anchors 5.5 m long were installed in the roofs of the large chambers to stabilize the excavation (Fig. 20-44), and gunite and wire mesh were used where the rock surface was highly fractured.

358

Fig. 20-44. Excavation and support of exhaust valve chamber of the Norad Expansion Project [170]

Chapter 21 S T A B I L I Z A T I O N OF R O C K A N D S O I L S L O P E S BY A N C H O R I N G

The stability of a rock or soil slope depends on its gradient and height, the stresses (vertical and horizontal) within the slope, the weight and strength of the ground as it has been naturally formed, the pressure of water in the soil pores or rock joints, and the effect of various external forces, such as permanent and changing loads impinging on the surface, shocks of all kinds, changes of temperature, etc. Slope surfaces are formed over long periods of time by the activity of all these factors, as well as atmospheric processes and sometimes also the effects of vegetation which break up the surface and reduce the stability of the slope. The main forces contributing to slope failure are those arising from the dead weight of the rock, soil, or other materials which place a load on the slope and from the pressure of water in the slope. These forces tend to bring about a downward movement of material as long as a gradient exists. A restraint on this tendency is provided by the shear strength of the ground. This strength is considerable where the ground consists of solid rocks, although it is often reduced by planes of discontinuity (cracks, joints, failures, fracture zones). It is substantially lower in the case of soils. The stability of slopes can be increased to good effect by anchoring them into the bedrock below the probable shear surface. The prestressing of the anchor increases the effect of friction at this surface and creates forces which directly act against possible movement%af the slope. The anchoring method can also be applied profitably when artificial slopes are created in the construction of roads etc., and in open pit mines. If the slope faces are secured with anchors as excavation proceeds, steeper gradients can be created. This not only means economy in terms of land, but also reduces the cost compared with earthmoving for unanchored, flatter gradients. This is shown by data given later in Chapter 29. Anchoring is most applicable where the superficial layers of existing or newly formed steep rock slopes need to be strengthened. A prestressed superficial layer of rock, unless it undergoes weathering, may well serve instead of concrete retaining walls, and at far less cost.

360 21.1 CALCULATION OF A N C H O R I N G FORCES

The stability of a slope is threatened by the tangential forces resulting from the weight of the rock or soil above the shear surface, any additional loading of the slope, and the pressure of percolating or retained water. Stability is maintained by cohesion and friction along the slide surface. The anchoring forces needed to stabilize a slope against shear failure are such, that when they are included in the system of all the forces acting on the shear surface, the condition of equilibrium is fulfilled within the required safety margin. 21.1.1

Soil slopes

The surface along which failure most often occurs in soils is cylindrical. The equilibrium of such a surface is usually expressed according to the method of K. E. Petterson, as described in any textbook on soil mechanics. The stability of a slope is maintained when the moments of those forces acting on the cylindrical shear surface (i.e. moments contributing to slope stability) are greater than the moment tending to bring about slope failure (Fig. 21-1): Σ AN.f.r

+ Ic . Δ / . r ^ AT. r,

where AN = the normal component (with respect to the shear surface) of the weight, AG, of a vertical strip of ground (AN = AG. cos a), / = coefficient of friction of the soil ( / = tg φ), r = radius of shear surface, c = cohesion of the soil, Δ/ = width of a strip of the shear surface, AT = sum of the tangential forces acting on the strip of the shear surface.

77

^^^^P7^7P^^

Fig. 21-1. Effect of the dead weight of the soil and that of a prestressed anchor on the shear surface beneath a slope

361

Slope safety demands that the stabilizing (passive) moments should be greater than the moments tending to cause failure (active moments). This difference, or safety margin is given by the relation: ΣΑΝ./

+ ΣΑΤ

Ic.Al

A more detailed analysis of the stability of earth slopes is given by Q. Zäruba and V. Mencl [231]. If an anchoring force, P, acts at the shear surface, this force contributes to the stabilization of the slope by virtue of the normal component, P„, and the tangential component, Pt (see Fig. 21-1). The safety margin for an anchored slope is given by: ,

f(I AN + Pn) + Ic . Al ΣΑΤ -Pt

If the axis of the prestressed anchor is deflected from the perpendicular to the shear surface by an angle ψ, then, Pt = P . sin ψ and, Pn = P . cos ψ. The necessary prestressing P, of the anchors is given by: p = mlAT-flAN-Ic.Al 1 m . sin ψ + / . cos ψ

.

The optimum value for the angle ψ has been stated by Hobst (see Chapter 4) as being the complementary angle of the angle of friction, φ: tg Φ = -j = cotg φ. The importance both of the inclination of prestressed anchors, and the coefficient of friction along the shear surface, for the degree of prestressing needed to stabilize a slope is apparent from Fig. 21-2. The effect of inclining anchors is more pronounced when the friction at the slide surface is low. However, as the angle between the prestressed anchors and the normal to the shear surface increases, the effect diminishes. The attainment of slope stability depends solely on the magnitude and angle of inclination of the resultant of the anchoring forces (Fig. 21-3). The point at which the resultant of these forces intersects the shear surface is of no particular importance, unless the coefficient of friction changes along the shear surface. If the latter is the case, the anchoring forces should be

362

Fig. 21-2. Dependence of the required anchoring force on its angle of deflection from the perpendicular to the shear surface (deflected so as to oppose the active tangential forces), taking the components acting at the shear surface as N = T = 1 MN, with safety factor m--= 1.5

Fig. 21-3. Diagram of forces acting on strips of prestress-anchored soil above the shear surface of a slope. A uniform height of the strips is assumed a) — in the vicinity of the emergence of the cylindrical shear surface at the foot of the slope, £)__at the lowest level of the slide surface, c) — near the upper edge of the shear surface located in the zone of the maximum coefficient of friction. Usually it is convenient to arrange for the anchoring forces to act near the lower point of intersection between the shear surface and the ground surface. It is quite possible to obtain a favourable angle of inclination for the anchors in this

363

lower position, where the boreholes may be short and only slightly deflected from the vertical; the drilling of such boreholes is technically the simplest. The analysis of slope stability and the calculation of the necessary anchoring forces must be carried out assuming the least favourable shear surface. The position of the latter is usually not known, except in very simple cases. Thus the analysis involves a large number of similar calculations, which were formerly carried out graphically, but are nowadays done with the aid of a computer. For the setting up of the computer program, the method of taking slices is expedient, since not only heterogeneities of the ground and variations of pore water pressure, but also surface loads and the influences of the individual anchor forces can be taken into account. When the least favourable cylindrical shear surface is sought, the procedure described by Otta [160] may be adopted. First, a circular shear path is selected in the longitudinal cross-section of the slope, such that it passes through the soils of the lowest shear parameters found in the investigation (Fig. 21-4). The safety factor for this path (the sum of the stabilizing forces divided by the sum of the destabilizing forces, acting on the shear surface of unit width) is determined without considering any effect of anchors. Further centre points of sliding surfaces

-25

10.0 22.5

35.0

¥7.5

500

72.5

85.0

97.5

110.0 122.5 135.0

Fig. 21-4. Longitudinal slope cross-section, and data required for stability analysis according to Otta [160]

364 cylindrical surfaces with smaller and greater radii, but with the same centre locus are considered, and the respective degrees of safety are noted. Proceeding further, small changes are made in the locus of the centre of the shear surface (as shown, for example, in Fig. 21-4), until it is found that the calculated degree of safety cannot be decreased any further. The magnitude of the anchoring forces required for stabilizing the shear surface of the smallest degree of safety is then calculated, taking into account the required safety margin also, and the positioning and lengths of the anchors are determined. To conclude, the overall stability of the anchored slope is assessed by substituting the parameters of all the relevant forces into the computer program. The use of computers has made it possible to analyse the stability of slopes and retaining structures of all types (see Chapters 22 and 23) in situations far more complex than those that could be handled by graphic methods or simple calculation. Thus, for example, the program of the Swiss Stump Bohr Co. has a capacity for dealing with 15 soil layer boundaries, 10 surface loads, 5 horizontal loads, 3 ground water levels, 1 ground water table, 10 pore water isobars, 15 anchor lines and 5 lines of support structures (for example piles). 21.1.2

Rock slopes

The stability of a rock slope is usually threatened when there is a reduction in rock strength, and shearing or tension develops along one or more of the natural planes of discontinuity transversing the rock slope. The greatest danger is from fractures and joints running approximately parallel to the slope surface, and inclined away from the slope at an angle (a) which is smaller than that of the slope gradient, but greater than the possible angle of friction,
365

/ / Fig. 21-5. Simple stability analysis of a rock slope with plane, parallel shear surfaces inclined in the same direction as the slope. Diagrams of forces are shown for: a) — an unstable slope (φ < α), b) — a slope at the limit of equilibrium (

°» V

oc

til

Fig. 21-6. Securing of a stable slope (φ > α), threatened by collapse along a plane slip surface, when loaded by an additional force, Q

of being upset by external forces such as additional loading of the slope (Fig. 21-6), ground water pressure, or atmospheric effects, the required prestressing of anchors acting perpendicularly to the slip surface is computed from the basic formula: P„ =

m T

-l.c f

N.

The forces acting normal to the slip surface are increased most efficiently

366

when the anchoring forces also act perpendicularly to this surface. However, structures can be secured more effectively against shear failure when the inclination of the anchoring forces to the slip surface is less than 90°, as pointed out in the preceding Section (see Figs. 21-5 and 21-6). Although the pressure, Pn, on the slip surface is reduced, a force component, Pt, is created acting parallel to the tangential destabilizing force, T, but in the opposite direction. The greatest effect of anchoring forces is achieved when they are deflected (by an angle, ψ, equal to 90° — φ) from the normal to the slip surface, or when their angle of inclination to the slip surface is equal to the angle of friction along this surface (as can be seen from the graphical solutions in Figs. 21-5 and 21-6). The value required for an anchoring force deflected from the perpendicular to the slip surface by an angle ψ, is usually computed from the formula: p> =

T

- l . c - ^m - . N

f

sin φ + -*— . cos \1/ m where T N / / c

= = = = =

sum of tangential forces acting along the shear plane, sum of forces acting normal to the shear plane, tg a = coefficient of friction along the shear plane, length of the shear plane, cohesion of the plastic filling of the shearing joint, or the effect of unevenness (indentation) of a strong rock along the shear plane on the joint shear strength.

If reliable data from field tests are not available, c is assumed to be zero for hard rocks, or the effects of cohesion or unevenness are included in the value for the angle of internal friction, φ. Usually it is difficult, if not impossible, to obtain precise values for φ, since otherwise the condition of limit equilibrium of the slope in its natural state could be expressed by: tg

or

φ = a,

where a is the gradient of the slip surface. The degree of stability could then be increased by anchorage, etc. to obtain the required value for w, e.g. m = = 1.2.Usually, however, it is more economical and generally more reliable to ascertain the angle of friction for the simple sliding of at least one small rock block over another. Z. Roth [181] analysed a more general case of slope stability with one or more joint systems inclined away from the slope face. He came to the conclusion that where there is a definite slip surface, a rock slope is permanently stable, regardless of the frictional resistance, provided that the slope

367

intersects this joint slip surface along its line of steepest descent (pitch line) (Fig. 21-7). If ß is the maximum gradient of a stable slope, a is the gradient of the planes of the joint system, and ζ is the angle between the slope face foot line and joint system, respectively, then

Fig. 21-7. Three-dimensional diagram of the structure of a permanently stable rock face with one system of planes of discontinuity inclined towards the face (according to Z. Roth) Ρι,Ρζ,Ρζ — planes of discontinuity of the joint system, a — gradient of joint planes, s — dip vectors of joint planes, ξ — angle formed between joint planes and rock base, β — gradient of the permanently stable slope. The shear surfaces of labile rock blocks in the vertical face of the cutting are shown hatched

Of all the joint systems that are surveyed in the field, those which are considered as being not conducive to stability have a gradient angle flatter than their probable angle of friction, φ. For the remainder of the systems, the permanently stable (slope) gradient is ascertained by substituting the angle of the least favourable case into the above formula, that is, the smallest angle β computed for an individual joint system inclined in the same direction as the slope. If the proposed or existing slope is steeper, the weight of the rock above the plane of the permanently stable gradient must be secured by anchoring. The simplified calculation of the required anchoring force (incorporating a greater safety margin) can be done by assuming a slip surface of inclinatin /?, parallel to the slope, and a substitute angle of friction φη > φ — j8, as in the preceding case in Fig. 21-5. The total anchoring force, P, allocated to a unit width of slope, is divided equally between as many individual anchors up and down the length of the slope as are needed to ensure that all potential slip surfaces above the foot of the slope are intersected and locked. If necessary, shorter local anchors can also be used (see Figs. 21-5 and 6-3).

368

The above-described solution to the stability of slopes threatened by planes of discontinuity inclined away from the slope and in a different direction from that of the slope, is very simple but very uneconomical. One may often see rock slopes with a V-shaped failure zone, brought about by two surfaces of discontinuity inclined in different directions (Fig. 21-8). For this reason some authors (Talobre, Peck, Goodman, Londe, Wittke, John, Hoek), have considered the possibility of a three-dimensional solution to the stability of wedge-shaped masses on rock slopes. A graphic method was worked out using the projection of a hemisphere on a plane (Fig. 21-9), on which three-dimensional diagrams of forces could be constructed. Using slope crest line of intersection

slope face rock wedge

a)

Fig. 21-8. Pictorial a) and natural b) view of a rock wedge failure

369

great circle representing slope face

Pensioned anchor average friction angle
Fig. 21-9. Simple stability analysis of a slope using a stereonet [91]. The slope is potentially unstable when great circles (representing planes of discontinuity) intersect each other in the shaded region (the plunge of the line of intersection exceeds angle of friction, i.e., %pf > ψι > > ψ)

Fig. 21-10. Optimum anchor positioning for the reinforcement of a rock wedge [91]

this projection, the planes representing the surfaces of discontinuity within the slope appear as circular arcs and the force vectors as straight lines passing through the centre of the hemisphere. The angle of friction, φ, of the shear surfaces (the cohesion c is neglected for the sake of obtaining greater safety) is represented by a circle with a radius proportional to φ (see Fig. 21-9). After having marked out the various parameters, it is easily shown that the rock wedge is unstable on account of its weight, when the intersection of the shear plane circles falls within the hatched area delimited by the circle of the slope face and that of the angle of friction, ψ. The method makes possible the solution of much more complicated cases of stability. For example, the shear surfaces delimiting a labile rock wedge may have different angles of friction, and the wedge may be subjected to forces other than its dead weight, such as water pressure in the joints, the dynamic effects of shocks, loads on the slope surface, and of course the forces exerted by anchors installed to give the required degree of stability [91, 64, 100]. At the present time, however, these problems are mostly solved analytically with the help of small calculators or computer programs. The method of calculation is very clearly explained, for example, by J. W. Bray [91]. A simple rule for the anchoring of rock wedges emerges from all these solutions: the tensioned anchor should be

370

aligned according to the line of intersection of the two shear planes, and it should make an angle equal to the average angle of friction with the line of intersection (Fig. 21-10). In rock slopes and faces in which the dominant planes of discontinuity are dipping towards the slope, the equilibrium analysis and determination of the necessary anchoring forces is also complicated. A completely reliable transfer of forces to the bottom region of the slope, across a joint system of angle of inclination, a, is only possible when the angle between the plane of the slope and the perpendicular to the planes of discontinuity is equal or less than φ (Fig. 21-11). The stable part of the slope is delimited by the

Fig. 21-11. Anchoring of a rock face with planes of discontinuity inclined towards the slope. The labile part of the rock formation is shown hatched; Kx to K3 are anchors placed in various directions in relation to the joints

theoretical stability plane (of inclination 90 — α + φ) which passes through the slope foot. In the slope above this plane the rock beds experience shear forces at the bedding joints, or tension from the bending of undamaged beds. The problem may be solved graphically and numerically, using the methods of various authors. Another approach to the investigation of the equilibrium of a rock slope with this orientation of the discontinuity surfaces has also been introduced [143, 225]. However, all the methods referred to here are rather laborious, and recourse is often taken to making the safe assumption that a limit slip surface forms in the plane of the steepest stable slope. The assumed labile rock mass is still delimited by a bedding joint which appears at the surface at a distance of 0.1 to 0.5/7 from the crest of the slope; this joint opens on failure of the slope. The anchoring forces required for any

371

selected gradient (Kl, K2, K3) are determined from the force diagram, and the total force is again divided equally among the individual anchors. More complicated slope stability situations in which the surfaces of discontinuity are parallel to the slope and dipping towards it or vertical, can be successfully solved by the well knownfiniteelement method [64]. A suitable computer program must be available for such an analysis. By this method it is possible, for a slope of given shape, to locate zones of tensile stress which are always sites of potential failure. In thefield,open joints may be observed at such sites. A typical example of the use of this method was given by Bukovansky and Pierce [27]; this concerned the stability of a high cutting made in a rock slope composed of a horizontal sequence of marlstones and shales with vertical joint planes running parallel to the direction of the slope. Fig. 21-12 shows the calculated lines of the principal stresses in a diagrammatic section of the slope, both before excavation, and after excavation of the step-shaped cutting of overall gradient 3 to 1. An outstandingly tensile zone appears, prior to excavation, behind the upper edge of the slope. After excavation, the zone is still affecting the rock ledges, but the stresses are of much smaller value (max. 0.2 ksf = 10 kPa). Tensile stresses of this magnitude can be handled effectively with horizontal rock bolts.

tensile stresses

0 6 12 18 meters negative, stresses in KSF (1 KSF=50kPa appro*.)

Fig. 21-12. Model of the slope in Parachute Creek Valley, showing rock joints, and minor principal stress contours before and after excavation [27]

372

A relatively simple method of computing the anchoring forces required to secure rock slopes in open-cast mines and quarries has been published by a group of specialists, K. Barron, D. F. Coates and M. Gyenge of the Canadian Mining Research Centre in Ottawa [8]; included is a favourable cost analysis of the method (see Chapter 29). The authors begin with the fact that where there is a complicated, dense network of natural discontinuity planes in a rock mass it is extremely difficult to determine in advance the most probable slip surface (unless there are outstanding stress zones and fault surfaces). In the method, a hypothetical slip plane is considered, this being inclined away from the slope face in the same direction, but at a different angle, and passing through the slope foot (Fig. 21-13). The angle of this plane

v=4-(cotgß'Cofa<*)

Fig. 21-13. Diagram of a rock slope with a hypothetical shear plane at which the specific tangential force τ is at a maximum, V — volume of labile part of the slope of unit width, / — length of the shear plane

with the horizontal, β, is not known in advance, but may be determined from the required slope inclination, a, and the probable or measured coefficient of friction, / = tg φ, between the existing joint surfaces of the rocks, taking the condition of maximum active force. The cohesion, c, usually does not have a constant value at the discontinuity surfaces of the superficially loosened zone of a rock mass, and is simply neglected with respect to the slip surface in the interests of increasing the margin of safety. The part of the slope above the assumed slip surface will tend to collapse, whilst the part below is permanently stable. If it is assumed that the labile part of the slope is a rigid body able to slide integrally along the slide surface (as is often the case in hard rocks), then by a simple analysis of the equilibrium of forces on an inclined plane, the tangential force, T (which after substraction of the respective frictional force must be secured by opposing anchoring forces), can be ascertained. The value of this shear force with respect to unit area of the slide plane inclined at an angle β, is computed from the equation: yH 2 τ = —- (cotg β — cotg a) (sin β — tg φ . sin β . cos β), where y is the volume weight of the rock in the slope, and the other symbols are as indicated in Fig. 21-13. The value of τ is determined by the slope gradient, a, and the coefficient of friction,/, for constant values of y and H. For every value of a and φ, there will be a corresponding value for /?, for

373

which τ will be maximal. This maximum value can easily be determined by differentiating the equation for τ with respect to /?, and by ascertaining the value of β. The authors of this method thus computed all the values of β for a in the range 0 to 90°, and / = tg φ in the range 0 to 1, assuming that α ^ β. The results are shown graphically in Fig. 21-14. From this diagram, β can be determined for known values of a and φ. Thus, from the equation H P = m .τ sin β the anchoring forces giving a safety margin of m = 1.5 can be calculated for a strip of the slope of total height / / a n d unit width. The same computations can be applied where a higher slope is divided by stabilizing benches. In this case the angle a represents the mean gradient of the entire slope, as shown in Fig. 21-15. The effect of ground water on the stability of a rock slope must be taken into account where considerable gradients exist, and where water can accumulate in the joint system, there exerting a hydrostatic pressure on the major vertical faces. The horizontal component of this pressure can very substantially upset the equilibrium, even in a rock slope formed of large blocks [231]. A lesser, but nevertheless continuous effect of water is its action on the rock surface, causing mechanical and chemical weathering.

50

60

70

slope angle oc

Fig. 21-14. Relationship by which the inclination, β, of the shear plane of maximum shear stress, τ, can be determined for a given slope gradient, a, and coefficient of friction, /, in the shear plane [8]

374

Fig. 21-15. Anchoring scheme for a stepped slope [8] pk — anchoring forces, K — cable anchors prestressed with force pk, n — reinforced concrete beams (sills) connecting the anchor heads, s — wire nets, u — fixing section of anchors (several metres long) in the stable part of the slope, π — theoretical conical zone of distribution of anchoring force in the rock mass

21.1.3

Dimensioning of non-prestressed anchors

Slopes should always be stabilized by using prestressed anchors. Where there is any danger, for whatever reason, that the prestressing might disappear, or in the case of small-scale operations in which the forces are small and therefore for the sake of simplicity the slope is locked to the bedrock with non-prestressed anchors, it is to be expected that prior to activation of the anchorage a partial displacement of the slope along the slip surface will occur. As a consequence of this the shear resistance of the rock is reduced, because the cohesion factor, c, is lost, and the angle of static friction is replaced by the angle of kinetic friction, φ Γ , which can be as much as 30 per cent, lower. The quantity of non-prestressed anchorage required to secure a slope with a curved potential slip surface must be that which safely transfers a tensile force, Pnr, perpendicularly across the slip surface, where: P- =

ml AT -ΣΑΝ, fr

or a force, Pr9 inclined with respect to the slip surface:

375

r

mIAT-frIAN m sin ψ -f fr cos ^

To secure a slope with a plane potential slip surface, the non-prestressed anchors must be such as to transmit a perpendicularly oriented tensile force, P'nr, given by: nr

fr

'

or an inclined tensile force, P'r, where: p t = r

m.T-fr.N m . sin ψ + fr. cos ψ '

The actual forces Pnr9 Pr, P'nr, Pfr generally need to be twice as great as the calculated values when the prestressing is introduced in advance; hence, it is recommended that nonprestressed anchors be used for stabilizing small slides only. The coefficient of kinetic friction also comes into effect in soil slopes in the active zone near the upper edge of the slip surface. It is therefore profitable to introduce prestressing forces into this zone also, in order to prevent any initial movement of the ground with the consequent loss of cohesion, and transition from a situation of static friction to one of kinetic friction. When a rock slope is threatened with collapse along a plane surface, and where failure occurs, as it often does, at the slope foot, anchors should be concentrated at these sites, and should provide an adequate degree of locking of all the beds that are traversed by dangerous planes of discontinuity. In other cases, the distribution of anchors over the entire slope is not necessary either, because solid rocks are usually strong and are sufficiently incompressible to remove any possibility of their gradually slipping down from the upper edge of the slip surface, as occurs on soil slopes. In fact any movement tends to occur simultaneously over the entire surface, and it therefore suffices to secure the most dangerous region near the slope foot. L. Müller introduced an original method of slope stabilization (usually applied to rock) [143], in which a system of non-prestressed steel ropes is laid on the slope surface, either across the slope (with the contour lines), or down the slope (with the lines of steepest gradient) according to which is the more efficient (Fig. 21-16). The ropes are anchored at both ends into concrete blocks, or directly into the stable strong rock. In the first case, where a slope bulge is anchored, the great advantage of the method is that there is minimal interference with the labile part of the slope, and yet a large force can be created. This arises from the fact that an arched anchoring arrangement provides a support (i.e. a force opposing

376

Fig. 21-16. Stabilization of a rock slope by laying a surface reinforcement of steel ropes [143] a) parallel with the contour lines, b) parallel with the lines of steepest gradient

movement) nearly twice as great as the tensile force in the ropes. In the second case, in which the ropes are laid up and down the slope and are anchored into concrete blocks (sills) at the foot of the slope and above its upper edge, the main purpose is to relieve the rock at the foot of the slope which is under the greatest stress, and transfer a part of the dead weight load to the upper, less heavily loaded part of the slope. In both cases the condition of the anchorage can be checked at any time since the entire system is accessible, and further ropes can be added if necessary. The stabilization of a rocky ridge slope with horizontal surface-laid steel ropes fixed to either side of the ridge, was carried out in the German Federal Republic. The densely fractured front slope of the basalt ridge was found to be sliding into the brown coal open-cast mine at Hessen. The slope was stabilized with four steel ropes of combined load-carrying capacity 2.4 MN (Fig. 21-17A). The lateral anchoring blocks were embedded in sound bedrock. Another, much larger scheme of this type, in which foundations were stabilized under historically important buildings, was carried out in Kassel [47]. Stabilization with horizontal anchored ropes provides a useful means of preserving structures generally.

377

Fig. 21-17. A — Sliding basalt ridge stabilized by means of horizontal, anchored steel ropes

CROSS -

F

SECTION

ig. 21-17. B — Design of high mountain slopes anchoring for coal pit mining in Czechoslovakia (according to J. Zajic)

378

Stabilisation of several hundred metres high mountain slopes of crystalline rock above very large pit mines in Tertiary brown-coal basin is under preparation in north-western Bohemia at present. The principal stabilizing system will consist of huge 200 MN anchors made of steel ropesfixedin 100 to 200 m long galleries (Fig. 21-17B). 21.2 STRUCTURAL ANCHORING METHODS

Slope movements and the destruction of slopes can occur over large areas along deep-situated shear surfaces. In rock slopes, it is often only the superficial parts of the slope, individual rock blocks, or the superficial rock layers damaged by weathering, that are in danger of collapse. The design and method of applying slope anchorage must therefore take into account the size of the anchored part of the slope and its importance. Reliable functioning of all the proposed components of the system must be ensured, and maximum effectiveness and economy must be kept in mind. 21.2.1 Stabilization of slopes The conditions for anchoring slopes to prevent them from slipping along shear surfaces extending through the entire slope or through a major part of the slope, were discussed in the preceding Sections of this Chapter. Besides the magnitude of the anchoring forces, the positioning of the anchors on the slope is also important. The anchoring forces calculated to be necessary, are usually divided equally among the individual anchors, or among lines of anchors running up and down, or across the slope. On non-uniform slopes, the anchors are placed on ledges, at the feet of the steeper parts of the slope (see Fig. 21-15). The loosened rock on the slope surface between individual anchor heads is secured by means of concrete sills with reinforcement and wire netting attached to the anchors. The dimensions of the latter structures can either be calculated by making a simplifying assumption concerning the effective rock weight [8], or, as is more frequently the case, they can be empirically designed according to the local conditions. The distribution structures, such as bearing plates, reinforced concrete sills, pillars or grids, are assembled from precast elements or are concreted in situ. The width of distribution sills and their foundation depth both principally depend on the type and quality of the rock. In strong hard rocks, the dimensions can be reduced to a minimum. In soils, on the other hand, bases 1 m2 or more in area have to be used. However, where soils are likely to

379

freeze, the bases must be founded sufficiently deep to prevent their being lifted by freezing, with a pressure greater than that of the prestressing. On rock slopes, the superficial rock layer on which the anchor heads bear may become weathered, and its compactness may be reduced. This surface must therefore be protected with gunite or reinforced concrete ribs, and those parts of the surface that are apt to fail must be covered with reinforced concrete slabs, or filled with concrete. As mentioned earlier, rock slopes are best stabilized with anchors sited in the lower part of the slope, particularly near its foot, with a suitable angle of inclination (Fig. 21-18). However, it is sometimes convenient, or even

o llol ollol o Mol o o o o ollol <^·Β·Κ=-

—=J —ZJ — —| ZZJ J z==-■^•■o

Fig. 21-18. Favourably placed concrete sills for the support of anchor heads at the slope foot

necessary where-cuttings are made, to place anchors first in the upper part of the slope, thus improving the stability as excavation of the cutting progresses. Inclined anchor boreholes are usually made in lines along different cutting levels (see e.g. Fig. 21-15). Sometimes, anchors are installed in vertical boreholes (Fig. 21-19), although this practice is statically incorrect since a vertical anchoring force always has a tangential component acting along the potential slip surface, thus increasing the danger of movement. This unfavourable effect can be partially eliminated by bracing beams, oriented up and down the slope [195], fitting between the transverse horizontal sills which support the anchor heads. Anchors installed on slopes are fixed into the mass beneath or behind the slip surface (or beneath that surface which delimits the labile rock body). The fixing depth measured from this surface must be such as to ensure sufficient rock resistance against extraction of the anchor (see Chapter 10). Due regard must be given to the possibility of other ground stresses developing, or the appearance of new slip surfaces immediately below the roots of the anchors. In any case the positioning and length of the anchors must prevent the development of any potential slip surfaces in the slope.

380 Fig. 21-19. Distribution of vertical anchors stabilizing a cutting for the Oakdale highway in California. Such anchors are only suitable for securing the bases of large rockslide masses, the anchors being placed in large diameter boreholes

Rock prestressing brought about by anchoring is particularly useful where transverse tensile stresses need to be eliminated. These stresses are caused by the concentration of pressure under structures situated on the edges of rock formations, or they may exist, for example, in the fixings of arched dams into the sides of valleys (Figs. 21-20 and 21-37), or under the foundations of bridge piers in the vicinity of deep cuttings. In Mexico, the centre pier of a bridge which was founded on piles and subjected to horizontal forces, was secured in this way (Fig. 21-21).

1111

Ϋ/Λ*ψ4ΜVAS.WA*U*y/4*

Fig. 21-20. Relief of transverse tensile stress in a rock by means of prestressed anchorage: a) — where a structure is sited at the edge of a steep rock formation, b) — at the fixing points of arched dams

381

I !

I i

I i

I i

" "Π

O

Fig. 21-21. Stabilization of a bridge pier founded on piles cast in situ (Mexico) 1 — anchors, 2 — piles of the bridge pier, 3 — pier

W\\x*\\o\VA\\W<

oo 21.2.2

Securing of rock blocks

Rock slopes often contain detached blocks whicrTate liable either to slide downwards over a plane of discontinuity, or to collapse when the rock strength is overcome at a point of weakness. Stability of these blocks is achieved by locking them to the stable rock mass with anchors. The necessary anchoring forces are computed in the same way as those required Tor the stabilization of rock slopes threatened by slides along plane shear surfaces (Section 2L1.2). If the block seems likely to fall by tipping up, the anchorage design* must stake into account the magnitude of the component (moment) of the block dead weight giving rise to movement. The stabilization of rock blocks on slopes by anchoring may be usefully complemented by the insertion of concrete beneath the labile block, or by grouting the surrounding joint spaces. Concrete applied under the block forms a sill and has a marked stabilizing effect; the anchorage may be used to secure the sill thus formed (Fig. 21-22). On a steep slope, where a rock block may be displaced as a result of being overturned, the anchorage must pass through the centre of the threatened block, close to its centre of gravity, Smaller rock blocks are anchored with short non-prestressed bolts fixed into the rock with grout along their entire length; the grout also protects the

382

b)

Fig. 21-22. Securing of rock blocks overlooking a roadway near Passau [190] a) — front view, b) — cross-section; 1 — secured rock blocks, 2 — reinforced concrete sill, 3 — concrete filling, made in belts (/—IV), 4 — supporting concrete buttresses, 5 — cable anchors prestressed to 500 kN, c — view of actual construction (photo PolenskySc Zöllner)

bolts from corrosion. For larger rock formations, prestressed cable anchors may even be used. The anchor heads are sometimes located in a small hollow cut in the rock surface, and are covered to preserve the natural appearance of the rock. The region of Cretaceous, boulder-shaped sandstones in Northern Bohemia typically suffers from block-like disintegration of the rock on the upper parts of valley slopes. After a survey, a large programme of conservation was drawn up, mainly relying on anchoring techniques. In

383

Fig. 21-23, a scheme is shown for the stabilization of individual rock columns and blocks near Decin, where a large collapse of rock occurred several years ago.

Fig. 21-23. Scheme for securing an articulated sandstone face near Decin (Czechoslovakia) with a system of anchors (groundplan horizontal section)

27.2.5

Protection of the surfaces of rock slopes

Stable rock faces can be a source of falling rock fragments and pieces of oose rock. Once they become detached, these fragments accelerate and present a danger to structures, vehicles, or people nearby. Anchored nets provide an efficient and economical means of eliminating this danger, for although these do not strengthen the rock, they retain any fragments detached from its face. The minimum anchoring forces necessary for the nets are obtainable at a relatively small depth into the stable rock surface. Protective nets are usually made of wire mesh or welded wire grids of various wire diameters (3 to 6 mm, depending on the size of the fragments to be retained). If necessary, double nets can be used. The wire grid is supplied in rolls of varying mesh size. For less exacting purposes, ordinary zinc-plated wire fence netting (3 mm diameter wire with a 5 x 5 cm mesh) is satisfactory (Fig. 21-24). When work is carried out on slopes and under rock faces, effective protection from falling fragments and pieces of rock is provided by synthetic fibre nets (Figs. 21-25 and 14-13). These are very elastic, strong, non-corrosive, easily handled, and give good adherence to the rock even on a very uneven surface. However, they are liable to severance on the sharp edges of large rock fragments, and they are more expensive than wire netting. Lengths 2 to 4 m wide, are most convenient for this purpose. Such nets are

384

Fig. 21-24. Protection of a rock slope near Freiburg with wire nets (documentation /. Kaim, Wien)

used only temporarily, and are usually fastened near upper edge of the slope, otherwise lying freely on the slope face. Wire nets are anchored to the rock surface with short non-prestressed steel bolts in boreholes 30 to 80 cm deep, according to the rock quality. The bolts are regularly distributed on the slope at a maximum density of one bolt per 4 m 2 of the rock face. The bolts are of the deformed bar type, with a minimum diameter of 20 mm, and a thread at the external end. They are inserted into boreholes 34 to 38 mm in diameter and grouted with a thin grout along their entire length. When the grout is hard, the protecting net is tightened against the rock surface with the nuts and larger washers of the bolts. The service life of these wire nets is estimated to be 15 to 20 years.

21.3 EXAMPLES OF SLOPE STABILIZATION BY ANCHORING

21.3.1

Stabilization using prestressed anchors

The first application of prestressed anchors as a means of stabilizing the slope of a cutting in cohesive soil took place in Czechoslovakia in 1961, when a section of the railway cutting near Cebin was saved. The cables used in this instance consisted of 7 wires of 4.5 mm diameter. The lower ends were fixed into anchoring cavities made by blasting, and the upper ends were

385

Fig. 21-25. Rock slope above a railway line near Prague covered with synthetic fibre nets. These served as a temporary protection for workers during the drilling and installation of anchors

fixed into Horel anchor heads supported on prefabricated load distribution slabs 80 cm square. The anchors were spaced 170 cm apart, and were each prestressed to 120 kN (Fig. 21-26). The successful saving of this section justified the proposed measures and showed them to be cost-effective, even though the project was not completed as regards one very important aspect of the structural detail, namely the foundations for the load distribution slabs. The slabs were simply laid on the slope face, which was of a soil type susceptible to swelling when frozen. Thus the anchorage experienced greater stress in winter when the slabs were pressed upwards by the frozen soil; and the anchor roots, which were pulled by a greater force than that which

386

^^φ: Fig. 21-26. Cross-section of the slope of a railway cutting near Cebin stabilized by means of prestressed anchors (VU1S 1961) 1 — anchor head, 2 — load distribution plate, 3 — anchor tendon, 4 — anchor root cavity, 5 — slip surface

they were designed to withstand, were partially displaced. In summer, when the soil was drier, the stress in the cables dropped. It appeared therefore that prestressing was only imparted from the anchors to the soil in the winter and spring, fortunately when it was most needed to ensure slope stability. This state of affairs is consistent with the fact that the secured experimental section of the slope has so far remained stable, while neighbouring sections have failed and the slope gradients in these sections have had to be reduced. A system of individual anchored precast slabs is nowadays used frequently for stabilizing soil and rock slopes, especially in cuttings (see Figs. 5-5 and 6-4). The slabs cover only 30 to 50 per cent of the slope surface and enable to design them in very steep gradients (60 to 70°). An example of a more complicated stabilization scheme in which anchors were installed in a slope of unstable soil is shown in Fig. 21-27. The slope drops away from a road leading to the construction site of a new building near Lutry.

it.

I

1

HIT—-- -L·^ Fig. 21-27. Stabilization of a slope at Tallepied 8, Lutry with Äßi?Fanchors for the construction of a new building, a — eluvium, b — old landslide, c — dense lake sands, 1 — most dangerous theoretical slip surfaces, 2 — BBRV anchors (2.4 MN, / — 23.30 and 34 m), 3 — planned building, 4 — road

387

The prestressing of rock by means of anchors also gave good results in the stabilization of a cutting at one end of the Ruzbachy tunnel on the Orlov— Podolinec railway line. The slope is formed in flysch beds of interbedded sandstones and shales. When excavation of the cutting started, a slope failure occurred and the work had to be halted until the rock had been anchored to a depth of 11 to 13 m in the substratum, with cable anchors (Fig. 21-28). The

Fig. 21-28. Cross-section of a cutting of the Podolinec—Ruzbachy railway line which was saved from collapse by anchoring the rock 1 — cables of 1 MN load-bearing capacity, 2 — load-distributing precast slabs, 3 — calculated slip surface, 4 — position centre line of the railway track

slip surface was found (from the digging of test pits) to be situated at an average depth of 6 m below the ground surface. The coefficient of friction,/, along the slip plane was 0.2, and the cohesion, c, was 10 kPa. The slope was secured by 212 anchoring cables (each with a load-bearing capacity of 1 MN) concentrated in a belt near the top of the cutting (Fig. 21-29). Rock slopes, natural or excavated, are particularly amenable to stabilization by anchoring, because the strong rock on the surface of the slope forms a convenient distribution layer for the compression forces exerted by prestressed anchors. An example of a very simple anchoring arrangement on a rock slope above a railway line in Czechoslovakia is shown in Fig. 21-30. Bar anchors of diameter 36 mm and length 7.5 m and 9.5 m were fixed with grout into a strong rock (phonolite), and then were prestressed to 100 kN and grouted in along

388

Fig. 21-29. Flysch beds of a slope in which the cutting of the Podolinec—Ruzbachy railway line was excavated

their remaining free length. A somewhat more complicated procedure than this was adopted to save a steep rock slope under the Orlik Castle which was threatened by predicted variations in the water level of the Vltava river in the basin of the Orlik Dam (Fig. 21-31). Bar bolts (30 mm in diameter and 6 m long) were placed in boreholes and grouted, to secure the protective cover of reinforced concrete to the dressed surface of the rock. The cover was concreted behind a stone lining, and had recesses for the anchor heads. The holes were sealed with masonry after the anchors had been prestressed to approximately 150kN and grouted to give a final smooth surface to the lining. The largest operation of this type being carried out in Czechoslovakia is the stabilization of a rock face 90 m high and more than 250 m long in the Labe valley, in the town of Decin [229]. The wall is formed of Turonian thick-bedded sandstones with a prevalent kaolinitic cement. Most of this cement was washed out on to the rock surface along well-marked bedding and vertical joints., Weathering of the bedding joints to a depth of 1 m from the rock surface has led to the formation of overhangs affecting the stability of the rock wall. Transverse and longitudinal vertical fractures of tectonic origin have also been gradually widened by freeze-and-thaw, which has even caused local displacements of rock fragments and boulders in these opened fractures. Owing to these destructive influences, separate blocks and pillars threatened with collapse rim the periphery of the rock wall.

389

'S

m

ÖL

%i

VVVV \

i Γ Fig. 21-30. Simple stabilization of a rock slope above a railway line near Usti n. L. (Czechoslovakia) by means of anchors and protecting wire nets

The static analysis was carried out for typical cross-sections where the slope was steeper or the overhang greater. The most frequently obtained stable angle was about 65° from the horizontal. The weight of the labile rock above a theoretical shear surface inclined at 65° (β = ψ) was simply determined from

390 Fig. 21-31. Anchoring of the rock face under the Orlik Castle 1 — granodiorite, 2 — veinous syenite porphyry, 3 — failure zone, 4 — depth of superficial weathering, a — injection boreholes, b — protective reinforced concrete lining anchored into rock

Fig. 21-32. Typical cross-section of the rock face near Decin, showing the arrangement of stabilizing anchors 1 — theoretical shear surface, 2 — design parameters of anchors (prestressing, length, spacing), 3 — exploratory boreholes situated behind the upper edge of the face in cross-sections Nos. 30 and 38; open joints are marked (4), as well as places of strong weathering of the sandstone (5), 6 — railway tunnel

Sec

-No.33

151m

nom 130 m

6 —■—*-* 0

5

10

15 m

0.6tlN\\ ά mnmfn—Λ 70/77 3 m

132 m

391

the area of an equivalent regular body of unit width and given volume weight. The equilibrium of forces on this shear surface, taking a safety factor of 1.2, may be expressed by the equation 1.2 G. sin 6 5 ° = G . cos 65° . tg 65° + Κύτι 65°. tg 65° + K. cos 65°, where K represents the total of the horizontal anchoring forces for the entire height of the rock wall. This yields the relation K = 0.768 G (kN). The anchoring forces required in the vicinity of cross-section No. 33, for example, (Fig. 21-32) amount to 1.04 MN. This total force was distributed over a number of anchoring levels in such a way as to provide the force needed at the foot of the slope, and to secure labile rock blocks at the surface, the stability of which was analysed separately. For these blocks, the degree of stability according to the ascertained state had to be assessed individually and the necessary anchoring forces had to be stated. The anchoring forces necessary for stabilizing individual blocks were determined using a safety margin of 2, because the parameters introduced into the calculation were taken from field investigations, and therefore did not incorporate a reserve factor as in the foregoing case of the theoretical slip plane. Following from the static analysis, several horizontal rows of prestressed cable anchors were designed. These anchors were 10 to 30 m long, with working loads of 0.2 to 0.6 MN. The entire rock face needed 436 such cable anchors, and 157 short bar anchors (Fig. 21-33). Effective anchoring of the rock abutments of the 233 m—high bowed Vajont Dam in Italy was the principal reason for the survival of the main dam structure without serious damage during the disastrous overflow of water in 1963. The overflow resulted from the sliding of the entire left-hand valley face above the dam into the basin. The slopes near the dam were strengthened with 300 anchoring cables 18 to 56 m long, prestressed to 0.5 to 1.0 MN. The support system for the anchor heads comprised vertical and horizontal beams and blocks, concreted on to the slopes (Fig. 21-34); the anchor heads were made permanently accessible by the installation of a system of climbing irons, gangways and catwalks. About 10 per cent, of the anchors were fitted with load sensors to allow the prestressing to be checked, and these were connected to the building on the top of the dam housing the controls for the entire dam. The disastrous overflow (a wave 100 m high was formed) as a consequence of the landslide swept away the control building together with the railings and all the access ways to the anchors, but the abutting anchored structures remained almost completely intact, as did the dam itself. In this case the large-scale anchoring of the valley sides was carried out by Polensky and Zöllner of the GFR. Fig. 21-35 shows a similar example of the anchoring of a high rock face at the bowed La Soledad Dam in Mexico, using anchors of the same PZ

392 Fig. 21-33. Securing of the rock face at Decin (Czechoslovakia performed by TSD and Stavebni geologie n. p.); general view (a), anchors installed in the face (b)

system (length 45 m, prestressing 1.4 MN). The often very simple arrangements by means of which complicated drilling, placing and prestressing operations are carried out for long anchoring cables, are particularly noteworthy (Fig. 21-36). At the site of the Kawamata Dam in Japan, prestressed anchors were used to secure a wall to the rock, the purpose of the wall being to distribute the pressure of the bowed dam into deeper, sound rock beds [56]. The superficial

393

Fig. 21-34. Steep rock slope below the Vajont Dam in Italy anchored by Polensky & Zöllner

rock strata were permeated by fractures and failure zones, and could not be relied upon to take the stresses transmitted by the dam unless special measures were adopted. The load-distributing cross wall was constructed along an extension of the centre-line of the dam, projected from the fixing point (Figs. 21-37 and 6-5). The cross wall was secured to the surrounding rock with prestressed anchors, using either the Dywidag type (total 155 MN) or the BBRVtype (total 85 MN) according to local conditions. The Dywidag anchors were made from high quality steel bars (diameter 27 mm), each prestressed to 0.4 MN (i.e. to 80 per cent, of the yield limit). Each anchor consisted of six parts interconnected by special couplings these parts being lowered sepa-

Fig. 21-35. Stabilization of rock face below the bowed La Soledad Dam in Mexico (photo Polensky & Zöllner) a) — view on the right-hand valley slope with PZ anchors, b) —transport of a 1.5 MN anchor to the site

395

Fig. 21-36. Anchoring of the steep rock face under the foundations of the abutment for the Rio del Oro bridge in Mexico, using VSL anchors 10 m long, prestressed to 400 kN

rately into the boreholes from a special gallery excavated for the construction of the cross wall. The prestressing of the tendons was adjusted after a month or two, and finally the boreholes were filled with grout along their entire lengths to establish an anticorrosive protection. The BBRVanchors were used to lock the foot of the retaining wall to the slope mass. At the site of the Aldeadavia Dam, the corroded beds on the left-hand side of the dam were stabilized with prestressed anchors before the slope was strengthened by pressure grouting; the designers feared that the pressure of the grout might loosen and lift the superficial beds on the slope. 169 cable anchors of 30 mm diameter and lengths between 12 and 36 m were used, and each was prestressed to 200 kN [150]. The anchors used to secure blocks against shear failure at the left-hand

396

I

0

1

5

1

1

1

10 15 20 m

C

B

A

^

Fig. 21-38. Stabilization of the bedrock of the left bank of the Santa Eulalia arch dam in Spain, by means of anchors A — quartzite bank, B — alternative quartzite and schists, C — siliceous schists

side of the 74 m-high Santa Eulalia arch dam in Spain, were designed to increase the strength of the bedrock under the dam by bracing the relatively narrow beds of shales and quartzites (Fig. 21-38). The anchors were 30 to

397

40 m long, and were fixed into sound quartzite beds unaffected by the excavations for the foundations of the dam. More than 1,000 anchors with load-bearing capacities of 1.0 to 2.3 MN were installed in the immediate vicinity of the heel of the 134m-high El Atazar arch dam. The anchors were needed to strengthen the slope, formed of much jointed shales and sandstones, at the left-hand side of the dam, which was threatened by the high water level. The anchoring heads were seated on a reinforced concrete grid placed on the slope surface (Fig. 21-39).

0 10 2030W5060 m Fig, 21-39. Slope reinforcement in the vicinity of the El Atazar arch dam (Spain), using a system of anchored reinforced concrete sills / — the dam body, 2 — anchors 1.0 to 2.3 MN, 3 — reinforced concrete grid, 4 — concrete slabs

The slope at the dam heel was protected by a continuous reinforced concrete slab on the surface. Nearly 3,000 zinc anodes were used for the anticorrosive protection of the anchors. The abutment of the Libby Dam, Montana, USA, was secured by means of prestressed anchors following the fall of a wedge-shaped mass of rock of 300,000 m 3 volume from the left-hand slope, during the excavation for the abutment foundations. On the basis of detailed investigation, the rock slope was secured with 50 VSL anchors (Fig. 21-40), each with a working force of 1.8 MN (60 % of the ultimate strength) and a length between 20 and 45 m. The fixing length was 6 m in all cases. The stability of the slope was continu-

398

Fig. 21-40. Anchored rock slope of the Libby Dam (photo VSL Corp.)

ously monitored by means of 5 measuring anchors. Since the anchors could not be visited in winter, the load sensors were connected to a central reading station. Securing the sides of construction pits by anchoring the exposed, and excavation-damaged labile beds of rock to sounder beds deep in the ground, is often considered as excessive additional work. This is all the more so where pits are excavated only for the short time needed to lay the foundations of a building. The work involved is generally very difficult, requiring highly skilled workmen and special machines; there are additional expenses which the client and the contractor are reluctant to accept, and so in order to reduce the construction costs and speed up the progress of construction, they take upon themselves the risk of failure of the pit slopes, and do not anchor them. The need for economy in the design and execution of engineering work undoubtedly involves a certain amount of risk, particularly where the structural arrangement and stability both depend on local geological conditions, as in the case of large construction pits. Generally, it is impossible to collect data on the characteristics of the ground in which the excavation is to be carried out with sufficient reliability to be able to make valid decisions about gradients and security measures for pit slopes. It is, however, necessary

399

to be able to determine the extent of the accepted risk; in an attempt to gain a relatively small saving by dispensing with anchors, far-reaching damage may occur, and additional outlay is then needed for repairs or for belated stabilization of the slopes, this outlay always being greater than that required for anchoring an initially undamaged slope. Moreover, a pit landslide will probably bring work to a halt, and delay the completion of the contract, which usually means further large financial loss. When a relatively long pit was excavated for the foundations of penstocks for the hydraulic power station of the Dalesice Dam in Czechoslovakia, the extent of the anchoring was limited, and the installation of the anchors was delayed in order to reduce the volume of construction work and speed up the progress of the operation. The result was a large landslide (Fig. 21-41 a). a)

Fig. 21-41. Landslide from a valley slope into a construction pit of the lOOm-high rockfill Dalesice Dam (Czechoslovakia) a) — view of the damaged slope, b) — see page 400

400

Fig. 21-41. b) — stabilization of the slope with anchors

The excavated valley slope contained a serpentine wedge between two blocks of damaged amphibolite. The slope had been stabilized by the installation of a line of anchors at about its midheight, but the necessary number of anchors and their dimensions were underestimated to such an extent that they were unable to prevent the landslide from occurring when the foot of the slope was cut away. The inadequate dimensions of the anchors were demonstrated by the fact that the anchors were severed in the landslide zone, at a distance of 40 cm below the anchoring head. The anchors outside the landslide region remained intact, and probably helped to limit the extent of the disaster. A great amount of unscheduled work was needed to prevent further earth movements. First of all, a loading rockfill bench was constructed at the foot of the slope, which was then relieved as the concreting of the penstock blocks progressed. The weight of the bench was substituted by several lines of slanting and vertical anchors fixed into the rock which remained unaffected by the slide at the bottom of the pit. Prior to the removal of the loading bench the slope was further secured by a line of anchors each of 1.0 MN, placed one third of the way up the slope (Fig. 21-41 b). The drilling of boreholes in the damaged rock of the landslide was

401 extremely difficult. The drilling proceeded from one level to the next down the slope, with simultaneous grouting or cement mortar filling of the boreholes, which otherwise would have collapsed. The grouting of the boreholes was complicated, by the presence of running water and the leaking of the grout through the open joints. Here, as in many similar cases, it was clearly shown how the coat of effective anchoring installed at the proper time in the slopes of the excavated pit would have amounted to only a fraction of the cost of salvaging the slope after a landslide. The disproportion is yet more marked when the time needed to carry out the work is taken into account in each case. The very strong and very little technically damaged rock masses of Scandinavia make possible not only the excavation of large underground caverns, but also the formation of other structures out of rock. Very large, stable, vertical rock faces can be formed, the surface being secured only by prestressed anchors. In Finland, for example, several docks 380 x 56 m, or 250 x80 m, with water depths of 9.5 and 12 m, were excavated in these rocks (Fig. 21-42). ¥.0f

□ DCüs

r" 1

•

/ <

A

"12.00

ii

Fig. 21-42. Anchoring of the rock faces of a dry dock in Finland (Oy Wärtsilä Ab's Perno). The anchors are Dywidag bar anchors of 32 mm diameter protected with steel sheet ducts, lengths and spacing as follows: 1 — 12 m, 2.5 m, 2 — 12 m, 5 m, 3 — (2 x dia. 32 mm), 12 m, 6.6 m

The stabilization of slopes finds application in many types of construction work. Rock prestressing made it possible to run a highway tunnel through a low rock headland, on which were the ruins of historic fortifications, in a narrow gorge of the Lueg Pass in the valley of the river Salzach in Austria. It was feared that the upper ridge of this headland might collapse whilst the tunnel was being driven, and this part was therefore anchored first with a row of

402

30 m-long cables (1 MN). These were fixed into the massive part of the slope behind the headland; only then did the cutting of the tunnel commence (Fig. 21-43). Smaller blocks in this section were secured with separate anchors installed on the slope according to need.

Fig. 21-43. Securing of a rock headland in the Lueg Pass in the Salzach Valley (Austria) with movable anchors PZ of 0.5 to 1.0 MN prestressing

The stability of galleries built below steep, high mountain slopes to protect highways from falling stones and avalanches, can also be achieved by means of anchoring. This method was used, for example, in Switzerland on the Zurich — Sargans highway along the steep banks of the Walensee lake (Fig. 21-44) or in Austria (Fig. 21-45). Another example of a purpose-designed anchorage system is that which was used to stabilize the foot of a cutting for the A 8 motorway in France. The foot of the cutting was strengthened with a system of concrete beams, oriented in the direction of maximum slope, and anchored at three levels. The prestressing of the anchors was 540 kN in the highest line, increasing to 1,080 kN in the lowest line. Precast reinforced concrete slabs were slipped into rectangular grooves in the sides of the beams (Fig. 21-46). With this arrangement, the progress of the construction was speeded up and an aesthetically pleasing view of the slope was created.

403

b) c) Fig. 21-44. Anchoring gallery foundations into the foot of a rock slope near the Walensee lake (Switzerland) a) — transport of VSL 1.22 MN anchor cables to the slope foot, b)—inserting the cables into the boreholes, c) — bottom part of the gallery with two rows of anchors

21.3.2

Stabilization of slopes with non-prestressed anchors

In practice, slopes exposed to shear stresses are sometimes secured with non-prestressed, or insufficiently stressed anchors. Such a solution is, however, uneconomical, because the installation of a large number of anchors means a larger amount of boreholes. Moreover, anchors that are exposed to direct shearing do not guarantee complete safety, least of all where the rock is soft, because failure can be caused not only by the shearing of the tendons, but also by the pressing of the tendons into the anchoring rock. This method of slope

404

Fig. 21-45. Securing of a motorway cutting in Austria by means of a retaining arch wall concreted and anchored in steps from the top down (documentation "Bureau BBR Ltd.)

section ΑΆ

s

,^Τ^Λ y

Fig. 21-46. Stabilization of the foot of a cutting for the A 8 Esteret-Cöte d'Azur motorway in France 1 — anchored ribs, 2 — retaining slabs, 3 — porous concrete, 4 — anchors, 5 — motorway stabilization can only be used in a modified form in which the shear-stressed anchors are located in concrete-filled boreholes or in pits of sufficient diameter for the rock face to rest on the concrete wrapping of the anchors.

405

An example of such a solution is given by a variant of the technique that was used to save a railway bridge founded on a slope; the slope was expected to become unstable, since its foot was to be inundated by the rising water behind a dam still under construction (Fig. 21-47). The slope was prevented from sliding by the installation of concrete blocks, which cut across beds of sandy and clayey loam and other horizons through which the slip surface

Fig. 21-48. Securing of an unstable slope of the Presna Dam (Poland) 1 — bedrock surface, 2 — anchors

Fig. 21-47. Stabilization of a slope by means of anchored concrete blocks under shear stress 1 — prestressed concrete blocks, 2 — cable anchored in bedrock and prestressed to 2 MN, 3 — shaft filled after the prestressing of the block

was expected to pass. The concrete blocks reached down to the bedrock and were anchored to it by means of prestressed cable anchors. Extra safety against shear failure under the slope and the retaining wall was obtained by increasing the size and prestressing of the blocks. Such a stabilization method is technically simple, because it involves a gradual, step-by-step excavation of the pits, concreting of the blocks, and prestressing of the blocks by anchors fixed into the bedrock. The slope of the Presna Dam lake in Poland was stabilized in a similar manner (Fig. 21-48). Non-prestressed anchors should always be placed so as to intersect the shear surface at the most acute angle possible, since shearing movement is then resisted by a larger cross-section of the anchor, and when movement begins, tensile stress is created in the non-prestressed anchor and becomes effective at the earliest possible moment. This method has been used, for example, in Jurassic stratified limestones to secure a rock cliff above the fixing zone of the left abutment of the bowed Chaudanne Dam, on the river

406

Verdon in France (Fig. 21-49). The anchor cables were composed of 125 wires of 5 mm-diameter high quality heat-treated steel with a strength of 1,400 MPa. The wires were placed in 13 layers separated by spacing nets to facilitate complete envelopment in grout. The diameter of the cable was 90 mm and that of the borehole 100 mm. The rock in the vicinity of the boreholes was grouted thoroughly from the top downward, whilst the cables were grouted in the opposite direction [221].

Fig. 21-49. Anchoring cables securing a rock cliff on the left bank of the Chaudanne Dam (France) 1 — rock mass threatened by collapse, 2 — critical slip surfaces, 3 — anchors ► Fig. 21-50. Securing of the rock slopes of the Kukuan Dam (Taiwan) 1 — beds of quartzite and sandstone, 2 — shafts (1.8 x 1.5 m), 3 — dam abutment

Short shafts filled with concrete have been used for the stabilization of rock slopes; the concrete filling is connected with concrete facing slabs on the slope surface by means of steel cables. This method was used for securing the slopes of the 94 m-high Kukuan arch dam in the central mountain range of Taiwan, completed in 1961. The rock slope, composed of beds of quartzite and sandstone, was to have been stabilized by prestressed anchorage, but since the Chinese engineers were highly skilled at tunnelling, they decided on a system of reinforced concrete prisms (1.80x 1.50 m in cross-section); these were made by filling specially made shafts (drifts) with concrete (Fig. 21-50). The shafts were driven at right angles to the slip planes (bedding planes), and were 6 to 12 m deep. According to the data cited in the report [147], this solution was the most convenient one under the existing conditions. The stabilization of rock slopes with steel ropes placed on the slope surface (see Fig 21-17) is also based on the principle of non-prestressed reinforcement, which is stressed by tension only after a slight movement of the slope has taken place.

Chapter 22 A N C H O R I N G OF WALLED

EXCAVATIONS

The walls of foundation pits and trenches up to 10 m wide are usually braced, but if they are wider than this, or it is necessary to have the maximum free space for the operation of earth moving machines and other construction equipment, the walls have to be tied back into the ground (Fig. 22-1). Anchoring is generally used where a deep wall is put in position in advance,

Fig. 22-1. Anchoring of the concrete walls of a foundation pit, thus allowing fully mechanized excavation (Langnau, Switzerland)

408

such as driven sheet piling, an underground concrete wall, lines of closely placed piles, or laggings held by pre-positioned soldier beams. The fixing of these structures into the ground well below the pit bottom is generally either unprofitable or impossible. The most suitable type of wall sheeting is selected according to the geological conditions, and this is anchored in stages from the top downward as the excavation proceeds, at levels determined from the static analysis. The first part of the calculation establishes the soil or rock pressures acting on the supporting structure; the second part concerns the anchor design, while the third part is an assessment of stability. 22.1 EARTH AND ROCK PRESSURES ON RETAINING WALLS The earth pressure on a retaining wall depends on the depth of the excavation, the physical and mechanical properties of the ground, and the deformation (angular or horizontal displacement) of the supporting structure. According to the extent of this angular displacement, the potential loading of the structure can range from the lowest active earth pressure, when the structure is deflected away from the earth, to the maximum resistance of the earth when the structure is pressed up against the earth, the load varying according to the degree of activation of the shear strength at the soil shear surfaces or rock discontinuity planes. If there is no deformation of the supporting structure, and the shear strength of the ground has not come into effect, the structure is subjected to a pressure at rest. Research on models carried out initially by K. Terzaghi [203] and then subsequently by other authors, has shown that a horizontal displacement of the top of the supporting structure of 0.001 to 0.005 of its height is necessary if the earth pressure is to drop to the value of active pressure, while an opposite and larger deformation is required to obtain passive pressure values. The necessary deformation for a favourable lowering of the earth pressure to its stress-activated value is obtained by slightly loosening the anchors at one level. The loading of a supporting structure which is anchored progressively from one level to the next manifests itself in a slightly different way. 22.7.7

Earth pressure

The earth pressure on a supporting structure is usually expressed in terms of a horizontal stress from the vertical stress in the soil, as determined by the soil volume weight, γ, the depth, A, and the respective coefficient, K: Gh = σν. AT, where σν = y . h.

409

The pressure increases in proportion to the depth, and can be represented by a simple triangular or trapezoidal loading diagram over the entire height of the support; the resultant acts at the centre of gravity of the diagram. The active pressure and the passive resistance represent the extreme values of the earth pressure with respect to the movement of the supporting structure and the soil. The equilibrium pressure settles at some intermediate value, since the actual earth pressure can, according to circumstances, assume any value between the two extremes. Pressure at rest This is the pressure exerted when no further deformation takes place. The coefficient of the pressure at rest is given by K0 = -

y

, or for a loose

soil, K0 = 1 — sin φ. In this equation, v is the Poisson ratio for the soil, and φ is the effective angle of internal friction. The total pressure at rest, S0, on the vertical rear face of a supporting structure of height //, with a horizontal ground surface behind the support, is: S0 =

^yH2.K0.

The resultant of the pressure acts horizontally in the lower third of the height, H, at the centre of gravity of the loading triangle. If another uniform load, q, rests on the ground surface, the earth pressure is increased by Δσ = = q . K09 and the resultant pressure on the back of the support is increased by Δ5 0 = q . K0 . H. The total pressure at rest is given by the sum S0 + AS0, and acts at the centre of gravity of the loading trapezoid. The pressure at rest in partly consolidated, or lcose, saturated soils is increased by the pore stress, or hydrostatic pressure. Extra'-pressure-^afr rest is also exerted on the supporting structure by swelling cohesive soils or by excessive mechanical compaction of the soil behind the supporting structure. Active pressure This pressure is exerted by non-cohesive soils only when the deflection of the upper edge of the support attains 0.001 H to 0.002//(for cohesive soils, 0.003// to 0.005//). The coefficient of this pressure, Ka, depends on the inclination of the supporting structure, the inclination of the ground surface behind the support, and the friction, cr adhesion of the soil to the rear of the support. Generally, Kah determined separately for cohesive and non-cohesive soils, according to current textbooks on soil mechanics. The design requirements usually involve a horizontal ground surface and a vertical support structure; friction on the rear of the support usually need not be considered.

410 For non-cohesive soils,

K. = t g * ( 4 5 - f ) , and the pressure of the activated earth, σα, at depth /?, is given by,

σα = ΊΛ.Κα

2

=

1ΛΑζ

\Α5-ξ\.

The total pressure acting on a support of height H, is again determined by a load triangle (Fig. 22-2), and the horizontal resultant, Sa, acting on the lower third of the height //, is given by,

Sa = l y . f i 2 . t g 2 ( 4 5 - f ) . surface of ground

"7

I 7(H-hc)Ka\

Fig. 22-2. Loading of a vertical wall by earth pressure. The ^ diagram is valid for loose soils, taking the loading height, H; it is valid for cohesive soils, taking the loading height H— hc

If the ground surface is loaded with a uniform, vertically-acting, sustained load, q, this is represented in the calculation by a layer of soil of depth H' = — , which increases the earth pressure on the support structure. y A similar approach is adopted in calculating the pressure of an earth body composed of beds of different properties. The pressure is determined for each bed separately, and the weight of the overburden is considered as a sustained uniform load, which increases the effective height of the bed. The loading of a supporting structure below ground water level is taken as the pressure of the earth in its state of buoyancy in the water, plus the hydrostatic pressure. According to the type of support structure used, and the soil type, account must also be taken of the frictional resistance developed at the rear of the structure, with an angle of friction of the soil of 0° to -γ-φ°1 this resistance

411

deflects the resultant of the pressure Sa upwards from the perpendicular to the retaining structure. For cohesive soils the pressure of activated earth at depth A, under the same simple conditions assuming a vertical support and horizontal ground surface is given by, aa =

y.h.Ka-2Cy[K;9

where

and c is the effective cohesion of the soil. The first component in this equation represents the pressure of loose soil with a uniform angle of internal friction, ψ. The second component represents the effect of the cohesion, c, which reduces this pressure. It is a characteristic of cohesive soils that they are able to maintain a vertical face of a particular height for some time, without exerting any pressure on the supporting structure. The critical height, hc, for this condition can thus be found from the relation σα — 0. By retaining the preceding equation we obtain,

The total active pressure of a cohesive soil on a supporting structure of height H, is given by, Sa = ^y{H-hc)2Ag2(^5

=

^j.

In this way the calculation of the pressure of a cohesive soil can be carried out as the calculation of the pressure of a loose soil of reduced height (see Fig. 22-2); similar graphic or calculation methods may be used to solve more complicated problems. If a cohesive and relatively impermeable soil has to be supported below ground water level, the hydrostatic pressure need not be considered, providing that the soil acts directly on the rear of the sheeting or wall. The pressure of a stratified medium is ascertained by taking each bed separately, and considering the overburden in each case as a permanent uniform load. The frictional resistance of the soil on the rear of the sheeting is estimated, according to the conditions, in terms of a deflection of the resultant total pressure, Sa, from the horizontal by an angle, 5, not exceeding the angle of friction of the soil, φ.

412 Passive resistance This is the pressure that is exerted where a supporting structure is pressed hard against the soil (e.g. at its lower, embedded margin) within the limit of shear failure of the soil. It is determined principally from a consideration of curved shear surfaces by using graphical methods; these methods can also be found in current text books on soil mechanics. For approximate calculations, or in situations where the support has a smooth rear face and the shear resistance at the plane of contact between the soil and the support can be neglected, plane shear surfaces may be assumed, thus simplifying the numerical solution. When the surface of the ground is horizontal and the rear of the support is vertical, the coefficient of passive resistance of the soil is given by,

K„ = tg*(45 + f-). For non-cohesive soils, the passive resistance is, at depth h

ap^y.h.Kp

= y.h.tS2^45

+ ^y

The total resistance of the soil on a structure of height, H, can be represented by a loading triangle, and the resultant, acting in the lower third of the height, H, is given by,

S,-^.*".tf(«+*). With a uniform, vertical, sustained load on the ground surface, the nonstressed earth pressure is increased by Ασρ = q. Kp. For cohesive soils, the assumption of a plane shear surface is too far removed from reality to be acceptable, and therefore a cylindrical shear surface is usually assumed for an approximate solution. Nevertheless, a simplified relation derived for loose soils is frequently used (as in the case of active pressure), when the passive resistance is increased by a permanent cohesion effect. The pressure at depth h, is given by,

σρ = γ . h . Kp + 2cV^7= V . h . tg2^45 + y ) + 2c tg^45 + | \ The total pressure on a structure of height, H, is obtained from a load trapezoid with sides equal to (2c y/Kp), and (y . H. Kp + 2c yfKp) respectively; the resultant acting at the centre of gravity of the trapezoid is given by,

413

Pressure on retaining walls anchored progressively at different levels. The pressure of earth on a retaining wall anchored progressively at different levels is analysed in a different manner from that of the preceding cases concerning a yielding wall anchored at one level. This situation is considered to be similar to that of structures such as sheet piling or framing, which are braced in stages. The upper line of anchors which is installed first, is already stressed by the time the next layer line is put in. The distribution of pressure is not, therefore, triangular with the maximum pressure occurring at the foot of the wall, but rather the pressures are shifted upwards, as confirmed by experimental data [204]. For loose soils, a uniform pressure is usually considered to be exerted over the entire support height. If the friction between the soil and the wall is disregarded, then according to the Czechoslovak Standard quoted [242] (Fig. 22-3), σχ = 0.65y . H. Ka.

Fig. 22-3. Distribution of earth pressure on a progressively supported retaining structure a) — non-cohesive soil, b) — cohesive soft soil, c) — cohesive stiff soil, or consolidated soil and soft rock

With a vertical support and horizontal ground, the coefficient of active pressure is given by Ka = tg 2 ί 45 - -~ 1. Hence, 2

ffl=0.65y.ff^g

i45-|\

This rectangular pressure distribution behind the support is 30 per cent. greater than the active pressure of triangular distribution. In soft, plastic cohesive soils (diagram b, Fig. 22-3), the earth pressure,. σ 2 , is given by, σ2 = y . H — Ac, whilst in stiff and consolidated cohesive soils and soft rock (load diagram c„ Fig. 22-3), the pressure, σ 3 , is given by, σ 3 = (0.2 to 0.4) y . H.

414

Some authors and also the compilers of the Swiss Standard [239] consider it correct to assume a uniform load of earth pressure over the entire height of the retaining wall (see Fig. 22-3a), when the latter is anchored at one or more levels in non-cohesive and cohesive soils. Therefore σα = 0.65γ. H. Ka. If a uniform load, q, on the ground surface behind the wall is considered, then, σα= 22.1.2

\.3(0.5γ. Η. Ka +

Ka.q).

Pressure of hard rocks on retaining walls

In excavation work generally, hard rocks are usually considered to be temporarily stable. If there are unfavourably oriented planes of discontinuity dipping into the excavation, either individually or in a group (a system of joints), then individual labile blocks, or the entire face are directly anchored without any other major supporting structures. This is carried out as described in Chapter 21.

22.2 DESIGN OF A N C H O R A G E FOR CASED EXCAVATIONS

In the design of the anchorage for a cased excavation, the dimensions of the foundation pit and support structure, and the geological characteristics of the ground are taken as a starting point. Usually a two-dimensional system of forces acting in the plane of a vertical cross-section of the excavation face is analysed statically; earth pressure and other forces are taken into account. The equilibrium of forces and the stability of the excavation both involve forces created by the prestressing of the anchor tendons which are secured to stable regions of the ground behind the support. In computing the earth pressure, the probable behaviour of the support structure when it is loaded must be considered. The values for the soil strength characteristics that are entered into the calculation are either reliable values obtained from field tests, or values taken from the Tables given in Chapter 9. The anchoring forces are distributed over the supporting structure with the maximum economy, that is, with the maximum negative moments at the anchoring points approximately equal to the maximum positive moments acting on the support between anchors. The prestressing of the anchor tendons is adjusted according to the assumptions made in the static analysis.

415

The anchors are fixed in a sufficiently stable region of the ground behind the supporting structure, and the overall stability of the entire structure is assessed. 22.2.1

Retaining walls suitable for anchoring

Types of retaining wall that can be anchored are sheet steel piling, soldier beams with laggings (Berlin method), piled sheeting, underground walls (Milanese method), and element walls (Swiss method). The static efficiency and distribution of earth pressure on the wall are influenced principally by the rigidity and continuity between components of the wall. Steel sheet piling is a relatively pliable, continuous supporting structure, which allows sufficient deformation above the excavation floor for the active earth pressure to develop. This characteristic is particularly valuable when the sheet piling yields along the line of the supporting anchors installed at one level. Sheet piles are usually driven below the excavation floor to a depth which still permits some angular displacement of the embedded pile ends to take place. The resistance of the soil to shear failure at the foot of the sheet piling prevents soil extrusion in the excavation floor. It is assumed in the static analysis that the displacement of the piles may even suffice to form a passive earth resistance. The loading diagram is shown in Fig. 22-4a. If it is expected that the displacement of the piling below the floor of the excavation will be less than that required to develop a passive resistance, a reduced value for this pressure must be introduced into the calculation, or if necessary, the

«r)

b)

Fig. 22-4. Earth pressure load on the anchored vertical sheeting of an excavation a) — loading of sheet piling, b) — loading of sheet bracing

416

part of the sheet piles embedded below the floor must be lengthened. A reduction of the earth pressure can be made, for example, by a reduction of the coefficient, Kp, by 1/3 of the value of Kp — K0, for the given soil. The Berlin method, using soldier beams with timber, concrete, or steel laggings inserted in succession, also represents a yielding type of sheeting (Fig. 22-5). Steel soldier beams are erected in, or sometimes rammed into,

Fig. 22-5. Sheeting comprising timber laggings (Berlin method), anchored in two rows (photo Brückner Grundbau)

vertical boreholes below the floor level of the excavation. Although they are very rigid relative to their small width, these beams still allow sufficient soil deformation for the development of an active earth pressure with a triangular load distribution. Each beam receives the earth pressure from a width of ground equivalent to the axial interval between beams. According to Krey [110], the passive resistance of the soil, Sp, acts to prevent any angular displacement of the beam around the point of a one-line anchorage, while the resistance, Rp, acts against shear failure of the soil on either side of the beam. This resistance is expressed in terms of the resistance

417

to movement along two parallel vertical planes FLM (Fig. 22-4b). In loose soils, Rp is determined by the friction at the vertical shear planes Rp = Es. tg φ. The horizontal pressure, Es, acting on these planes, is expressed as the pressure at rest of the given soil (with K0 = 1) on the area FLM. Thus,

£s = !v.
Rp = c.d\tg(45

+ ^\.

Piled sheeting comprises a row of piles, with intervening gaps greater than the pile diameters (Fig. 22-6). This sheeting method is suitable for cohesive soils or soft rock, which will be retained by the piles without sheeting owing to the arch-forming effect of their cohesiveness. The exposed surface may be strengthened with gunned concrete. Conditions in the gaps between very rigid piles are highly favourable for the development of active earth pressure. The displacement of the piles in the soil does not fulfil the excavation face

gunned concrete

I I t D' i t

earth pressure

t t i

Fi 22_6 Pile st ckade

s· ·

( <> )

sheeting (horizontal section)

conditions for an active earth pressure above the excavation floor, nor for a passive resistance below the floor. Therefore it is recommended, especially when there is no significant yielding of the anchors, that the coefficients Ka and Kp in the load diagram be reduced to the values K'a and Kp, respectively, as in the following: K'a

=

Ka + — (K0 — Ka),

K'p = Kp —— (Kp — K0), and that the calculation at the anchor level be carried out, taking the earth pressure at rest (Fig. 22-7).

418

it

|.

tfKj

|.

t_

r(Htd)KZ

j,

Fig. 22-7. Earth pressure and earth resistance loading on an anchored pile wall (K' values), and an underground wall {K" values)

Underground walls for supporting the soil are usually continuous reinforced concrete structures cast in deep, narrow trenches. This type of structure is very rigid and creates a soil angle cf friction of — φ to —φ at its rear face. It is recommended, therefore, that the coefficient of earth pressure and earth resistance be adjusted in the following way: 2 K"a = Ka + —{KQ — Ka), K'p = Kp — — (Kp — K0). This is shown in Fig. 22-7, where the earth pressure at rest must also be considered at the anchoring level. Element walls are made and anchored in a succession of horizontal levels from the top downwards, as the pit excavation proceeds (Fig. 22-8). Their advantage over the preceding methods is that any drilling, ramming, or excavation with special machines, does not have to be carried out prior to the opening of the pit. The individual reinforced elements are made on the spot and the reinforcements of the elements are welded together as the elements are formed.

419 front

view

cross-section v///////,

>///////// O

b

1

O

llll

1

° 1 ° I °~T I ° I

O

o

0

o

|l|l|l|l|l|l|l|l

v)//////}///////\ v///;///;///;//////;///, >///////;//;

x

/V// νλ>////, >////////)//

v///////;

° I ° I ° 1 W///////V))})///

y/////////

°x I * 1 ° I 1 °1 \ ° , \ ° \

v$/\))/\ e

'///////////////},

Ύ777777777

^7 >/////////)/,

Fig. 22-8. Wall of anchored concrete elements — construction sequence [158] (A), anchored element wall for the Beaux Arts underground station at Charleroi (Belgium), constructed by Bruckner Comp. (B)

Such element walls are yielding and the calculations therefore assume an earth pressure evenly distributed up and down the entire height of the wall. For the individual stages of the excavation, however, an additional arch loading on each line of mostly recently installed anchors arises from the

420

pressure of the yet unanchored next part of the excavation [158], and this has to be taken into account. For the bottom lines of anchors, this transient load will be greater than the sustained load (Fig. 22-9).

22.2.2

Calculation of anchoring forces

The required anchoring forces are calculated on the basis of the equilibrium of forces acting on the retaining wall. If the anchors are fixed in one line, the calculation is simpler than that needed for several lines of anchors one above the other. 22.2.2.1

Anchoring forces along one level of the retaining wall

A calculating method for one anchorage level is given by Z. Bazant [14]. The load-bearing wall is sunk below the floor of the excavation to a depth, d, the magnitude of which is determined by the requirement that the foot of the wall should not shear over the soil beneath it, nor rotate about the anchorage point K. The sum of the moments of all the forces acting on the wall with respect to the point K9 must be zero. Thus according to the scheme shown in Fig. 22-4b, *Vi + S2r2 - S3r3 - 2Rprp = 0,

421

where Sx, S2, S3 are horizontal components of the earth pressure or resistances already solved graphically, and Rp is the resistance to shearing along vertical planes on either side of the beam, as indicated previously. By substituting the respective dimensions of the structure and the soil parameters into the equation, values for the individual forces and their distances from point K are determined (according to the loading diagram) in terms of the unknown d. Substituting into the equation IMk = 0, a value of d is tentatively obtained from the biquadratic equation; this should then be increased by 10 to 20 per cent, (according to H. Blum [20]) for the actual construction of the wall. By entering the value of d, values for the individual forces are computed, with the exception of the anchoring force which is determined from the equilibrium of forces in the horizontal direction: Pk = St + S2 - S3 -

2Rp.

The dimensions of the retaining wall is computed from M m a x , which acts at the depth, y, (below ground level) at which P = 0; at this depth the earth pressure is exactly balanced by the anchoring force: Pk = Si. The steel anchor tendon, the abutting head, the load distribution beams YYi

P

and the anchor root are designed to take a force Pw = mPk, or Pw — —r—- 9 sina if the anchor is inclined at an angle a. The safety factor m in this case is determined according to an accepted Standard, or according to the data given in Section 17.1. The length lt of the tendon is initially taken as being equal to the excavation depth (lt = H) [50], and the angle of inclination is fixed between 10° and 30°, provided that the anchors reach the required load-bearing bed. The size and length of the anchor fixing depends on the type of root used, and follows the principles given in Chapter 13. In the final part of the design the overal stability of the anchored structure is assessed according to the principles set out in the next Section. The design must state the prestressing force of the anchors. It is recommended that some allowance be made for a small angular displacement of the supporting structure (1/1,000 H for non-cohesive soils, 3/1,000 H for cohesive soils), in order to satisfy the assumption that a stress-activated earth pressure will be exerted. This means that the anchor prestressing can be decreased in proportion to the displacement of the structure, and the stressactivated earth pressure on the supporting structure then adjusts the anchor prestressing to the computed value (see Chapter 23.1). Some experts have recommended, however, that even where a stress-activated earth pressure is assumed, the full prestressing force should be introduced into the anchors.

422

This approach must allow a margin of displacement in case a reduction of anchor prestressing takes place as a result of long term soil and anchorage creep. 22.2.2.2

Anchoring forces along several levels of the retaining wall

The supporting structure is treated as a continuous vertical beam supported at successive anchoring levels and at the extremity, F, of the embedded end; the beam is loaded by a system of earth pressures and resistances. The embedding depth, d, as the length of the lower part of the continuous beam, is determined first of all and is then used to determine the reactions of the continuous beam. The available passive earth resistance on the embedded part of the supporting structure, should be 50 per cent, greater than the basal reaction of the continuous beam at point F. If this safety factor of 1.5 cannot be fulfilled, the calculation must be repeated with another value for the embedding depth. Although this procedure takes much time and computation, the final correct analysis immediately gives the static values required for establishing the dimensions of both the structure and the anchors. The static analysis for the continuous beam, particularly in cases in which the system of forces is more complicated, may be carried out graphically, by transforming the distributed load into concentrated loads in horizontal lines, and applying the strip method. The embedding depth for a continuous sheet piling wall or an underground reinforced concrete wall (which, with different water levels on either side, also prevents bursting of the floor of the excavation) is not necessarily determined by the need to prevent shear failure below the excavation floor, but rather by the need to prevent bursting of the floor when the soil gives way under the pressure of the water. It is of particular importance that this requirement be satisfied where there is only a shallow embedding of the wall, this being sufficient to resist shear failure of the soil. The dimensions of the supporting structure and the anchors are established after the static values have been determined, in the same way as for anchors at one level. The tendon lengths of the lower lines of anchors may be reduced, but the prestressing of the anchors is kept at the full value. The static analysis is completed by assessing the stability of the anchored structure, as described in the following Section. 22.2.3

Stability assessment of anchored retaining walls

The security of individual anchors against extraction from the soil was discussed in Chapters 10 and 13. This security is achieved by using an anchor root (a plate or bulb) of appropriate dimensions at the fixing point of the

423

anchor in the soil. The system must be tested by loading test anchors before work begins on the anchored structure, and by overstressing standard service anchors with a higher, testing load (see Chapter 17). All this, however, is not enough to guarantee the stability of the anchored structure. It is also necessary to ensure that the tendons are of sufficient length and orientation to reach a region of the ground of which the load-bearing capacity is unaffected by the excavation. In principle, the anchor root is at a greater distance from the wall than that of the shear surface delimiting the soil wedge which loads the wall. Two shear zones are formed behind the wall, these being defined by an upper and a lower slip plane (see Fig. 22-11). The formation of two slip planes behind an anchored wall has been demonstrated in model tests carried out by H. Bendel [15] (Fig. 22-10). The fixing of anchors beyond the surface of the permanently stable slope, as discussed in Chapter 23 in the context of permanent retaining walls, is uneconomic for the temporary walls of excavations. To establish the most reliable anchor lengths and the conditions for of internal stability of the system: wall-earth body-anchors along the so called deeper shear surface, the methods proposed by E. Kranz [108] and N. Janbu [98] are most often used. Both methods were compared by Huder and Arnold [93]. cm

135-

^ ^ V,\ \ \ ' k\\\ \\ k I I Ä S\\\,u fr\ I ^\ vu ' \\\ \v w \\\ v < Λ \w\ \ \\\ I \ M \ I \ t\ \\ ^ *\ l 5 Λ \N\\ 1 1 \ \\ Si \ \ V ^ 1\ ' v \] s ^v|\ Λ > \ \ \ \' I I I

115

i

*v

I

35 75

'v

I

55 35 15

0

l

20

22.2.3.1

40

60

V

w w w

k

Θ0 100 120 Wem

Fig. 22-10. Failure of a sand block model with double anchorage; diagram shows motion vectors at the points of a grid in a perpendicular cross-section through the model [15]

Internal stability

The stability can fail when the soil mass behind the wall, compressed by the forces (Rk) of the anchors, shears along a plane connecting the theoretical foot, F, of the embedded load-bearing wall with either the centre of gravity of the mass, or the tips of the fixing devices, D, of the anchors (Fig. 22-11). In addition, when the anchor roots are placed near the ground surface, the soil may fail along a shear plane running through the points D, up to the

424

,

f t

*

^ X anchor plate (screen)

A6F\ loading CJDf wedges of soil FbCO earih body securing stability of anchoring area

Fig. 22-11. Stability analysis of a wall with different anchor fixing methods (after E. Kranz) / — anchoring by trench-wall or plate, / / — by long concrete root or tensile pile, / / / — by expanded root bulb

long anchor roof

anchor bulb

surface at point C. The stability solution according to Kranz is based on the equilibrium of all forces acting on the soil body CDFB, which ensures the stability of the anchorage behind the wall. These forces are: the weight of the soil, G1, the earth pressure, St, the resistances along the shear surface, Ra and Rt, and the limit force of the anchors, Rk. The latter is usually determined graphically from the diagram of forces. The diagram of forces can be constructed from the known vectors Gt, Sl9 Ra, from the direction, Rt, of the frictional pressure acting on the shear plane, and from the direction of the anchoring force, Rk. According to the usual practice adopted for a graphical solution, the distribution of forces on the slip surface is not taken into consideration. The calculation is simplified if the reaction Ra, which supports the labile soil wedge ABF, is replaced by opposite forces with which it is in equilibrium, viz. the weight of the wedge (Ga) and the reaction of the wall to the earth pressure (Sa). The force Ra is thus eliminated and the slip surface of the loading soil wedge ABF need not be

425

separately considered. It remains, then, to solve the equilibrium of the entire soil body CDFA, involving the weight of the body, G = Gt + Ga, and the forces Sa, Si9 Rt and Rk. The force diagram is considerably simplified in this way. Arithmetically Rk is expressed as follows:

_ A - A + [G - (hSa - A ) t g 3 ] . tg(y - δ)

R

*

cos a + sin a . tg (φ — #)

'

where hSa, hS1 are horizontal components of the relevant earth pressures (see Fig. 22-11), a = angle of inclination of the anchor, φ = angle of internal friction of the soil, $ = angle of inclination of shear surface FD, δ = deviation of earth pressure Sa and St from the perpendicular to the wall, as determined by the friction between the soil and the rear of the wall. It may be necessary to introduce the extra pressure of ground water into the static analysis. The sustained service load, q, of the ground surface behind the wall is introduced into the calculation only if the inclination of the slip plane FD is reduced to a safer, lower angle. The anchored structure is stable when the limit anchoring force, Rk, which disturbs the equilibrium at the shear surface DF, is 50 per cent, greater than the anchoring force, Pk, needed to secure the wall (see Section 22.2.2). The safety margin, m, is thus

If this condition is not fulfilled, the anchor tendons must be extended or inclined more steeply, and thus fixed deeper in the soil mass. A more precise solution for the stability of an anchored wall with a curved shear surface between D and F h a s been elaborated by R. Jelinek and H. Ostermayer [99]. They have demonstrated both theoretically and by experiment that in most cases of anchored supporting walls, shorter anchors could be used than those suggested by the approximate computation in which the lower shear surface is assumed to be planar. 22.2.3.2

Verification of internal stability under various anchoring conditions

The procedure just described was initially proposed for sheet piling anchored in a non-cohesive soil, the anchor tendons being fixed in the soil with the aid of vertical anchoring plates or trench walls (see Fig. 22-11, I). It was subsequently applied to other fixing methods involving a long or

426

expanded grouted root, or a rammed tensile pile (see Fig. 22-11, II, III). In the latter examples, a hypothetical vertical anchoring plane CD, is assumed for the purposes of the calculation. This plane passes through the centre of the anchor root or fixing section of the pile (length /„), or within 2 m of the expanded anchor root bulb (Fig. 22-11, III) [50]. When the anchors are arranged with their roots in a row, their spacing is assumed to be more than — /. For cohesive soils, the calculation is similar to that for non-cohesive soils but the earth pressure on the supports and the resistance at the shear surface FD are determined taking an effective angle of friction of the soil, φ', and an effective soil cohesion, c'. In unconsolidated cohesive soils, it is assumed that φ = 0. In the diagram of forces, a force C = c'. FD, acting along the slip plane, is introduced before the force Rt. When the soil is of more than one type, so that the soil mass behind the wall comprises two or more beds, of different physical and mechanical properties, the stability is calculated as follows (Fig. 22-12):

Fig. 22-12. Graphic analysis of the stability of an anchored wall in front of a stratified medium

The body AFDC is divided vertically into parts at the points of intersection between the shear surface FD and the individual beds. The weights of these parts always act in the diagram of forces from the points of intersection of Rly R2, etc. with the direction of the anchoring force, Rk. The section between the point of action of the first force, Sx, and the point of intersection of the previous component of shear resistance with the anchoring direction, determines the value of the limit anchoring force, Rk. If the supporting structure is anchored at several different levels, the stability of the structure is calculated in stages from the diagram of forces (Fig. 22-13). The state of equilibrium for the first line of anchoring forces is considered initially, (this involves the weight, Gl9 of the body FDXCXA and

427

Fig. 22-13. Analysis of the stability of a wall anchored at several levels in homogeneous non-cohesive soil

the forces acting at the shear surface FD^), and thus Rkl is found. The next force diagram is constructed for the soil body FD2C2A, the shear plane FD2, and the sum of the anchoring forces Rkl and Rk2. This procedure is repeated until the lowest shear surface is reached. The resulting safety factor is obtained from the relation

In Fig. 22-13, this analysis is shown for three lines of anchors in a homogenous non-cohesive soil. 22.2.3.3 External (overall) stability It appears from the results of research and from recent practical experience that an assessment of the external, or overall stability is necessary, particularly for high structures or where complex geological conditions exist. Here, the formation of a curved shear surface is assumed, the depth of this depending on the depth of the foot of the structure (Fig. 22-14). In the analysis, the least favourable shear surface with respect to the supporting structure is sought, using the above-described procedure for the internal stability. The position and shape of the least favourable shear surface is not known in advance and therefore the investigation requires a large number of similar calculations, for which a computer and suitable program must be available. To set up this program, it is convenient to use the method of slices as applied to the stability of a soil slope. The procedure for finding the least favourable shear surface may be that described in the preceding Chapter

428

Fig. 22-14. Overall stability analysis for a tied-back wall, assuming cylindrical slip surfaces. Geological parameters, the topography of the site, and an account of the loading are given according to L. Otta [160]

(21.1). The axis of the cylinder is gradually shifted as shown in Fig. 22-14, until it is found that the calculated degree of safety does not decrease any further. To obtain a more accurate result, it may also be useful to investigate the state of stability at that part of the shear surface where the cylindrical surface approaches ground level; the shear surface in this section takes the form of a plane inclined at the angle a = 45 ——, according to the theory of active earth pressure. If the smallest degree of safety found in this way is less than the required degree of safety (usually 1.5), longer anchor tendons must be used, or the number of anchors at the lower levels must be increased. The new design is assessed in the same way. Several hundreds of shear surfaces can be processed by the computer for one design of an anchored wall, and conditions can be introduced into the analysis that are for more complicated than any that could be handled by graphic methods. 22.3 EXAMPLES OF A N C H O R E D WALLS AND THE M O N I T O R I N G OF THEIR FUNCTION

A simple example of an anchored pile sheeting around a construction pit excavated in unconsolidated topsoil and soft clayey shales is shown in Fig. 22-15. The sheeting required to withstand an earth pressure of the

429

Fig. 22-15. Anchored sheeting of drilled-in piles in the foundation pit for the National Assembly Building in Prague

following parameters: φ = 35°, y = 2,000 kg/m 3 , δ = 2/3 φ (the angle of friction between the soft rock and the sheeting), consists of a row of drilled-in piles. These are embedded into the ground 3.5 m below the pit floor, and are supported higher up by an anchored reinforced concrete belt (anchors 1.8 m apart). The anchor cables are each composed of 20 patented 4.5 mm diameter wires. The roots, which are 4 m long, are fixed in shales. The pile sheeting of the excavation for an underground railway station in Prague was anchored by two lines of anchors mounted on longitudinal reinforced concrete load-distributing sills (Fig. 22-16A). Bar anchors with a load-carrying capacity of 370 kN (see Fig. 13-43) were fixed over a length of 6 m in weathered clayey shales, by repeated grouting with a collared pipe. An example of collapse in the anchored sheeting for the Prague underground railway is shown in Fig. 22-17. At the top of the excavation are about 10 m of compact loamy gravels and sands, beneath which lie highly weathered clayey shales. Braces, made of No. 40 steel I-beams, were inserted into boreholes extending below the pit floor. As the excavation progressed, wooden laggings were inserted between the beams and these were anchored successively by two lines of anchors which were slightly inclined from the horizontal. The first line was fixed 2 m below the ground surface in the gravels and sands, and the second was fixed at a depth of 12 m in the weathered shales. The sheeting was stable until the ground became over-saturated with water escaping from a burst water main situated below the ground surface 15 m from the sheeting. The pressure on the wall increased and the shear strength

430

Fig. 22-16. A — Anchored pile sheeting of the foundation pit for the Kacerov underground station, Praha,

of the anchored soil was reduced, both under the anchored soil mass and along the anchor roots. The soil failed along a curved surface emerging at ground level up to 8 m from the top of the wall; the anchors were torn out and the braces uprooted. Anchored drill-rod sets (Hagconsult system) have been successfully used in Sweden to anchor the sheet piling of foundation pits in dense urban areas. Fig. 22-18 shows a cross-section of the sheet piling for a foundation pit in Stockholm. The sheet piling is anchored by an upper line of bar anchors fixed in morainic gravels and sands, whilst the lower line is anchored in the bedrock. The bored anchor bars, which were of 32x16 mm cross-section, were grouted with cement slurry under a pressure of 0.5—2.0. MPa, and prestressed to 360 kN; they underwent elastic elongation only. A view of the sheeted pit is shown in Fig. 22-19. The external anchor heads were mounted on load-distributing rolled steel beams, and these in turn were seated on oblique props welded to the sheet piles. Anchored sheet pilings are also commonly used to support banks, and are used in the construction of embankments. Sheet piles are rammed into the underlying strata of the bank, and the ground level is raised up to the top

431

of the sheet piling which is then anchored with tie bars, usually positioned from above. The tie bars are connected at one end to the sheet piles, and at the other to embedded or rammed-in anchoring screens (or plates) (Fig. 22-20). Another example involving an embankment is given in Chapter 23. Anchoring is used in a similar way to secure quay walls constructed by the slurry trench method or the so-called pile-wall method. The quay wall at a port near Cremona in Italy may serve as an example; the port is built on a canal connecting Milan with the Adriatic Sea. The quay wall was constructed as an underground wall of reinforced concrete, 70 cm thick (Fig. 22-21). It was secured by tendons spaced at intervals of 3 m and prestressed to 540 kN. A parapet wall served as a base for the anchors; this wall was concreted in a trench 2 m deep below the ground surface, at a distance of 10 m from the quay wall.

432

Fig. 22-17. Collapse in the anchored wall of the foundation pit for the Prague—Budejovickä underground railway station

21.70 ^rayelandsand \

\ Qs^wage main

Fig. 22-18. Section through the anchored sheet piling for a foundation pit in Stockholm

The anchoring of a reinforced concrete wall at the sides of an excavation for a water canal leading from a nuclear power station in France, is shown in Fig. 22-22. Permanent anchors, fixed in sand, can be seen at three levels. The upper level comprises bar anchors with a load-bearing capacity of

433

tißr-

J?**'

Fig. 22-19. View of a foundation pit walled by sheet piling (Anchoring by Hagconsult AB Stockholm) t

embankment

sheet piling

Fig. 22-20. Detail of an embankment wall anchored with steel tie-bars

500 kN and a fixing length of 6 m while the two lower levels comprise rope anchors with a load-bearing capacity of 1 MN, and a fixing length of 13 m. The construction was carried out by the Soletanche Co. The underground walls of the pumping station supplying water to Bratislava (Czechoslovakia) provide an illustration of another method of anchoring. The station is sited on the flood plain of the Danube, where

434 3Ä30

Fig. 22-21. Quay wall at Cremona (Italy) / — reinforced concrete, 2 — anchors of 540 kN each, 3 — anchor base, 4 — anchor foot embedded in the concrete of the Milanese wall, 5 — anchor head on the base wall

Fig. 22-22. Anchored walls at the sides of an excavation for a canal in France (Soletanche Co.); ä) — view into the canal, b) — see page 435

both frequent floods and considerable fluctuations in the ground water level occur. The anchors had to be fixed in beds of fine-grained gravel and sand which are often saturated, and because of the small load-bearing capacity of anchors fixed in such material, an excessively large number of them would

435

'ft 2 r

^alfu'viaf-t. ~

vv

+7.5

"

Fig. 22-22. b) — cross-section

b)

ill i I

-±-

=^φψ TT I

I I I

JJ-

^ ^

**:

/ / /

! '' I

tfj

/

/

I / / I /

'

Fig. 22-23: Anchoring of the walls of the pumping station for the water supply of Bratislava (Czechoslovakia) with the aid of sunk open caissons a) — plan, b) — vertical section, / — underground wall of the pumping station, 2 — open caissons, J - 2 M N anchors, 4 — anchor heads

436

have been needed to secure the station walls. Thus it was proposed that the anchor bases might rest against the walls of open caissons lowered in advance around the circumference of the pumping station (Fig. 22-23). The anchors were to be made from 65 mm diameter single ropes with both ends fixed in cast steel heads. The design gave particular attention to anticorrosive protection, the individual strands of the anchor rope being coated with red lead during the stranding. The insulating wrapping of the rope consisted of a layer of insulating paste and an anticorrosive bandage, which was protected from mechanical injury during manipulation of the anchor by a PVC foil bandage.

horizontal

section

open pit

Fig. 22-24. Anchored element wall of the excavation for a new hospital at Chur (Svitzerland) a) — plan; b) — cross-section showing the anchoring system and the distribution of horizontal deformation in the wall [159]

When excavations are carried out in comparatively dry cohesive soils, the anchored sheeting is often constructed according to the Swiss method (see Section 22.2.1). An anchored element wall protecting the excavation of a new hospital at Chur [158] is shown in Fig. 22-24. The excavation was dug to a depth of about 16 m in heterogenous detritus below the old hospital. The wall was composed of reinforced concrete blocks measuring 3.25 x 1.90 m, and was anchored successively at eight levels with Stump-Duplex bar anchors 32 mm in diameter, with a working load of 450 kN. The movement of earth behind the wall was registered by inclinometers in two monitoring boreholes drilled before excavation commenced. The deformation is shown diagrammatically in Fig. 22-24b; the top of the wall was displaced 1.1 cm horizontally, and no vertical deformation was observed. The wall was pressed into the seil at about mid-height, most probably as a result of the greater anchor

437

prestressing that was applied, since the calculations were based on a wall load of 1.2 times the active earth pressure. The old building did not suffer any damage. An interesting example of the movement of an entire sheeting wall into the soil was reported by S. K. Saxena in the USA [73]. The wall, securing the excavation for the World Trade Centre building, was 18 m high and 1 m thick. It was constructed of reinforced concrete panels and was secured by six lines of anchors with a fixing length of about 10 m. The movements of 10 panels were observed throughout the construction. The wall acted as a semi-rigid member and tended to rotate about a point near its base, the position of this point being determined to some extent by the depth of the foot of the wall in the bedrock. It was observed that the wall moved continuously into the soil as excavation proceeded and successive layers of ties were prestressed. The maximum lateral movement was about 6.5 cm, recorded a year and a half after the excavation had been started. The anchor loads decreased steadily with increasing deformation of the wall, owing to elastic contraction of the prestressed tendons (Fig. 22-25). The excavation for a new telephone building in New York was protected by means of timber panels and steel battens anchored by long triple rope prestressed tendons grouted into the rock [73]. Above the centre section of the wall the tied-back concrete underpinning of the existing six-storey telephone building was realized. The total depth of the excavation was 23 m. A problem arose when the basement excavation was required to be taken down into a rock stratum underlying an earth overburden. The support system for the overburden could have been tied back with anchors dipping at 45 degrees into the rock, but these would have given rise to a vertical loading on the support system that would have been transmitted to the rock in which the support was resting, very close to the excavation face. This would have created a danger of rock failure at the face. Both the importance and the usefulness of monitoring excavation support structures are apparent from the example described by Otta [159]. An excavation 16.5 m deep, carried out in Zurich, was protected with a reinforced concrete underground wall anchored at four levels under unfavourable conditions (Fig. 22-26). The behaviour of the structure and the surrounding ground were observed during the progress of the excavation and afterwards by measuring the inclination of monitoring boreholes behind the wall, and by measuring the changes in the prestressing of the anchors. Systematic measurements commenced when the excavation reached a depth of 7 m, at the level of the second line of anchors (day 0, see Fig. 22-26). Until then, no deformation had been registered either at the ground surface or in the wall. The deformation was considerable, however, when the excavation reached the third anchoring level (after 16 days) and the fourth anchoring level

438

■^- IT|ti

. Ilil i»«:aaÄ :

^*>*<

rrpfi^

^^^^^mm^mm

600 (3.0j [

§ «50 νν. |(Ζ25)| \Υ^4ι

.^.V.-

-«««..---^:.~Γ.

1

-ϋ ΓΖϋ-·; " ν

.5- 300 (1.5)

%

·*Λ

- 1 .- _ ; :

^ - > "——~»

^Π?2 ^5

\ X

Αι 150 150

300

±50 600 750 time · days

900 1050 1200

ß Fig. 22-25. Anchors securing the excavation wall for the World Trade Centre [73] A — view into the foundation pit, B — measured decreasing loads in different anchors

(after 139 days). Gradually the ground behind the wall sunk by as much as 3 cm (at a distance of 5 m from the wall), the upper edge of the wall was displaced 4 cm horizontally, and at the level of the third line of anchors the

439 deformations and distances

cross- section

o 2 i 6 έ fe^fe1•L(control points observo^ — ^ ^tion borehole

Fig. 22-26. Measured deformation of the wall and displacement at the ground surface at construction pit in Zurich [159]

Λ*>.*A/ ^

horizontal displacement was 8 cm. On the basis of these measurements, the procedure by which the last stage of the excavation was carried out was modified in time to prevent the deformation exceeding the admissible value. The ground water level in the surroundings of the pit was lowered by pumping the water from drilled wells, and the last three-metre stage of the excavation was undertaken in parts, in each of which the concrete foundation slab was cast without delay. The increments in the ground and wall deformations were markedly reduced, both in the course of the last excavation stage and after completion of the excavation work. In the construction of the underground railway in Nuremberg, anchored sheeting was employed in a novel manner. The tunnels for the stations had to be excavated in sandstone at a shallow level below the surface, close to the foundations of the 75 m-high southern tower of the church of St. Lorenz. To avoid the risk of damaging this magnificent Gothic building, the tunnels were located on the opposite side of an anchored pile wall (Fig. 22-27). The drilled-in piles had a diameter of 80 cm, and the anchors were prestressed to 400 kN. During the excavation of the tunnels, deformation of the surrounding ground was carefully recorded, and it was found that the total vertical displacement of the foundations amounted to only 2.6 millimetres [12]. An interesting example of a braced and anchored inclined supporting wall is shown in Fig. 22-28 [222]. The wall was constructed to protect a foundation pit for the Munich subway tunnel and prevent subsidence under the foundations of a nearby multi-storeyed building. The wall was installed by means of a Benoto boring rig. The ground water level was not lowered. The stabilizing of excavations in plastic clays by means of anchored walls is very exacting, on account of the considerable pressures and deformations which occur in these soils. H. Breth and D. Stroh give an example of a large construction pit excavated in clay in Frankfurt a. M. [23]. The pit was 20 m

440 Fig. 22-27. The use of anchored piles to protect the foundations of a Gothic church against ground settlement in the course of excavation of the station tunnels for the Nuremberg Underground (GFR). a) — cross-section; b)—plan

securing of St Lorenz Church

a) cross - section Königstrosse L 9.60 , | * & , 9.60

i; \ ! / > j \

I

i / >

441

deep and 177 m long, and the computed force exerted by the earth per 1 lineal metre of the supporting wall was 1,900 kN. The excavation was secured by anchors installed at six levels in succession, beginning at the top, with a spacing of 1 m between anchors in the same line. The anchors were 20 to 27 m long, slightly inclined from the horizontal, and the injected cylindrical roots were 7 m long. The horizontal displacement of the top of the wall was regularly checked during the excavation (Fig. 22-29). Midway along the pit, the top of the wall was displaced horizontally by almost 12 cm in the course of seven months of excavation, after which time it remained stable. In view of this

Fig. 22-28. Inclined wall of a foundation pit in Munich / — street level, 2 — building foundations, 3 — subway tunnel, 4 — braces (H beams), 5 — prestressed anchors, 6 — ground water level, a — silty gravel, b — sandy silt, c — silty sand

f

**'

i

0 r-

(

/

f

120

— —

1

100 80 |

1

1 1

5■

|

1 1

/ (

10

40

1 J ^ J

15-

20

// XI

XII

1969

/

20

/ t 1

II

///

IV 1970

V

VI

VII

VIII

yo |

Fig. 22-29. Excavation progress (in terms of depth) and deformation of the wall of a construction pit at Frankfurt a. M. [23]

442

large deflection of the wall, the forces in the anchors were checked, and were found to correspond with the force of active earth pressure, which the anchors were designed to hold. In another example from the same locality [24] a large excavation in clay was secured by an anchored wall; in addition to the systematic observation of the loading and deformation of the wall as well as the anchor prestressing, comparative calculations of these values were carried out by the finite element method (Fig. 22-30). The results of the calculations and measurements were

calculated

measured

Fig. 22-30. Deformations and anchoring forces in the wall of a construction pit in Frankfurt (GFR), according to the results of both calculations and measurements [24] A — profile of the wall and geological profile; 1 — gravel, 2 — clay, 3 — limestone; B — deformation of the wall, C — earth pressure, D — increase in anchoring forces with excavation depth

in very close agreement, both for the deformations and the anchoring forces. Some differences appeared at the foot of the wall, where the effects of interlayers of strong limestone in the clay was not considered. The displacements of walls anchored in plastic cohesive soils, not only in the upper part but also at the foot of the wall, are a consequence of the compressibility of the soil below the excavation floor. The various forces involved are shown in Fig. 22-31 [154]. The earth block AB DE experiences a horizontal pressure, and in order to satisfy the condition of equilibrium, the resulting earth resistance at CD must be equal to the earth pressure acting on AE. This means that the soil below the floor of the excavation, compared with its original state before excavation began (pressure at rest) is now under a lower vertical pressure and a higher horizontal pressure, resulting in compression of the earth and displacement of the wall. In addition, shear strains are imposed on the block ABDE which must be compatible with the strains occurring below the excavationfloor.The displacement of the walls may be influenced favourably by the length of the anchors,

443

excavation

Fig. 22-31. Stresses and strains around an anchored soil block behind an excavation wall, according to Ostermayer

according to Breth and Stroh [23]; the magnitude of the displacement decreases almost linearly with increasing length of the anchors. On the other hand, the earth pressure and the anchor load are not affected by the length of the anchors.

Chapter 23 A N C H O R I N G OF SLOPE R E T A I N I N G WALLS

Retaining walls are usually built to take over the function (static effect) of a ground body removed from a natural slope. These vertical, or slightly inclined walls are constructed for the purpose of preparing a horizontal surfaces for roads (Fig. 23-1), water course (Fig. 23-2), or foundations for cranes, industrial buildings, warehouses, quays, and harbours. The purpose of the retaining wall does not usually have any bearing on its design.

Fig. 23-1. Anchored supporting gallery for the Lukmanier highway (Switzerland) 1 — slope detritus, 2 — 293 Tubfix anchors (137.5 kN)

Fig. 23-2. Anchored retaining wall for the covered part of a canal connecting the basins of the Tage and Tietar Dams (Spain) 7 — BBRV cables, 2 — strong diabase into which the anchors are fixed

The structure of the wall must be capable of transferring the lateral pressure of soils or rocks, together with the weight of the wall and any load placed above it, safely into the load-bearing foundation ground. The following basic conditions apply in the static analysis of a slope retaining wall. The force, N, normal to the bottom multiplied by the coefficient of frict i o n / , in ratio with the tangential force, T, acting at the foot of the wall must be equal to or greater than the safety margin for shear failure, ms (this is taken to be within the range 1.2 to 1.5). The ratio of the sum of the negative moments divided by the sum of the positive moments acting on the retaining wall must be equal to, or greater than, the safety margin for overturning, mp (this is taken to be within the range 1.5 to 2.0). The maximum stress in the footing of the wall must not exceed the safe stress of the foundation ground, which is ascertained on the basis of standard values or from load tests.

445

Anchoring the retaining wall into the underlying ground helps to satisfy these requirements efficiently. Slope retaining walls offer a wide range of possibilities for the positioning and orientation of anchors. The vectors of the anchoring forces can be vertical, inclined, or horizontal, and they can be arranged so as to pass through the centre of gravity of the cross-section of the wall, or even across its upper edge (according to which best contributes to the stability of the wall). When the stability of slope retaining walls is achieved by anchoring. objections are sometimes raised, unjustifiably, that the anchoring must inevitably lead to an increased pressure in the soil, since the wall is pressed against it and gives rise to a passive earth pressure as a result. Such a situation, however, can only develop if the moments or shear forces arising from the anchoring pressure, are larger than the sum of the forces arising from the earth pressure and the weight of the wall acting in the opposite direction. The application of anchor prestressing to this extent is obviously uneconomic and out of question, and a passive earth pressure as a reaction to the anchoring cannot be allowed to appear behind an anchored retaining wall. The greatest pressure exerted by the soil itself on the supporting structure is its pressure at rest. This occurs where the supporting structure, including the anchorage, is designed in such a way as to disallow even a small degree of deformation, the soil behind the retaining wall being held in its original position. When a linear or angular displacement takes place, the pressure of the soil gradually falls to the lowest value of active earth pressure. The displacement at the top of a retaining wall is assumed to range from 0.001 to 0.005 times its height, according to the soil type. A discussion of earth pressures on supporting structures was given in the preceding chapter. Slope retaining walls are comparatively massive, rigid structures, and their loading is therefore generally considered in terms of the pressure at rest, with coefficient K0 (see Section 22.1), or, if the active earth pressure is considered, the coefficient is increased by 2/3 of the difference K0 — Ka, to the value K"a (see Section 22.1.2). The load on the wall at depth, Λ, in a non-cohesive soil is determined from the expression: σΗ = γ.Η.Κ0

or

y . h . K"a

and for a cohesive soil: ah = y.h.K0

or

y.h.K"a-2cJK.

The additional loading of slope retaining walls by the hydrostatic pressure of ground water trapped behind them is reduced as much as possible by providing reliable drainage. Thus a vertical sand or gravel layer is interpo-

446

sed behind the wall, and this is drained from its base. In this case the effect of friction between the soil and the rear of the retaining wall need not be considered in the static analysis of the external forces.

23.1 CALCULATION OF A N C H O R I N G FORCES AND THE D E S I G N OF ANCHOR F I X I N G S

The value of the anchoring force required is determined simply from the equilibrium of the moments of forces with respect to the fulcrum (most often the external lower edge of the wall), if overturning of the wall is considered, or from the diagram of forces acting on the wall, if a displacement of the wall along its foundation is considered. The basic formulae for the static analysis were introduced in Chapter 3 and 4. Whether the problem is one of overturning or one of displacement, it is clear that there is an essential requirement for the foundation of the masonry to be sited on a stable and load-bearing rock bed. If active earth pressure is considered in the analysis, the necessary deformation of the support must be made possible. In most cases it is recommended that the anchors be designed so as to allow development of the full active pressure together with the necessary deformation of the wall; on average the latter is about 0.002//(see Section 22.1). The required stressing force, P'k, may thus be determined from the expression: P'k = , 5 ;

-XE.F.

If λ . / = 0.002//, then, P'k = &-0.002^-. where tS'a H λ / E F

= = = = = =

E.F,

force of the full active earth pressure on one anchor, height of the retaining wall, coefficient of ductility of the steel of the anchor tendon, length of the anchor tendon, modulus of elasticity of the steel, area of cross-section of the tendon.

where h is the safe stress of the steel.

447

Substituting for F, we have,

*-Ä(I-

0.0024^).

The partial yielding of the anchor head as a result of displacement of the anchor root in the soil, is neglected. The anchor roots must be fixed into permanently stable part of the ground behind the support. In a loose soil, the region below the plane sloping upward at the angle φ from the lower edge of the wall may be considered as being such a stable part. In cohesive soils, the stable region behind the support is located by identifying potential shear surfaces, using the well known methods of soil mechanics for the solution of slope stability such as Pettersson's strip method or the graphic solution of Fellenius, both of which are described in all good textbooks. For a rapid solution, the graph in Fig. 23-3 may also be used. With hard rock behind the retaining wall, the stable region is delimited by a plane inclined at an angle β; the determination of this angle where unfavourable planes of discontinuity are present was dealt with in Section 21.1.2. If hard rocks appear behind the retaining wall beneath a cover of soil, the stable region is usually taken as being within the rock mass. The anchor root must be fixed with its entire length within the stable region with due regard for the required safety margin against uprooting (see Chapters 10 and 13).

^tga Fig. 23-3. Graph for rapid determination of the limit gradient of a permanently stable slope in a cohesive soil (after Fellenius)

448 23.2 S T R U C T U R A L A R R A N G E M E N T OF A N C H O R E D SLOPE R E T A I N I N G WALLS

The loading of slope retaining walls generally remains constant, and therefore the centre of the anchoring forces should 'be located as far as possible from the centre of gravity of the footing of the wall, in the opposite direction to that of the loading pressure. Thus it is useful to increase the strength and width of the foundation by providing the structure with vertical buttresses (Fig. 23-4), by shaping the foundation like a cantilever (Fig. 23-5), or by bringing the horizontal force to act near the top of the structure

-

\

•

1

/;

Ί

\

J\1c 0 | . > «s

i J

< 3' '

/

I

1

560 660

100 *^ I "

660

Fig. 23-4. Example of the anchorage of a buttressed retaining wall loaded on one side 1 — foundation of a buttress, 2 — buttress, 3 — retaining screen

$

S2

Fig. 23-5. Loading of a cantilever retaining wall by both earth pressure (5Ί, S2) and the weight of supported material (Gi to G6)

Fig. 23-6. Retaining wall anchored by horizontal forces

449 (Fig. 23-6). The anchorage should be prestressed because this makes good use of the load-bearing capacity of the wall footing (Chapter 4), and is fully complemented by the masonry weight in contributing to the stability of the structure. Anchors which introduce horizontal forces only into the wall, do not increase the pressure in the wall footing, which is an important advantage where walls are founded on ground of low load-bearing capacity. On the other hand, horizontal boreholes are more difficult to drill than vertical boreholes, especially in non-cohesive soils and detritus. Retaining walls held by horizontal anchoring forces acting near the top are also loaded by bending moments, which they must be designed to withstand. The wall is secured against shear failure not only by the anchoring, but also by the passive resistance of the soil to extrusion above the footing at the wall face; the foundations are generally situated below ground level in order to protect the footing from frost. A wall anchored by horizontal forces is thus secured in the same way as anchored sheet piling or underground screens, except that in the case of a slope retaining wall the weight of the structure is an extra aid to stability. Anchoring finds application, both economically and technically, in all current types of retaining wall. Its economic advantage is increased, as already mentioned, in structures of reduced weight and increased stress moment (foundation width), such as buttressed walls and reinforced concrete cantilever retaining walls. In some instances it is possible to reduce the weight to such an extent that this factor can be entirely neglected in the static analysis; thus the retaining wall functions purely as a load-distribution plate supporting the anchor heads. 23.2.1

Precast slope retaining walls

In buttressed retaining walls, the load-bearing and stabilizing function is separate from the retaining function. These walls require less masonry and may be built from precast elements (see Fig. 23-4). Fig. 23-7 shows a retaining wall consisting of buttresses (cast in situ) which are stabilized by anchoring into the foundation rock; these support the retaining screens. The buttresses are positioned in individual rectangular foundation pits so as to prevent any possible loss of stability of the slope detritus, which would occur if continuous trenches were to be excavated for the foundation sills of a compact and uninterrupted retaining wall. By excavating relatively small and well braced foundation pits for independent buttresses, the stability of the slope is unaffected. Also once the buttresses are constructed and anchored, the slope can be cut away with earth-moving machinery to make a space for the foundations of a building. The stability

450

additional load of the slope by a building

Fig. 23-7. Buttressed retaining wall 1 — buttresses, 2 — prestressed cables, 3 — retaining screens (precast), 4 — labile bed of detritus, 5 — natural arch in soil

of the slope as a whole will be ensured by the buttresses and the natural soil arches which are formed in the detritus and which rest on the buttresses. The precast elements of the retaining screen only take the pressure of the detritus inside the natural arches and protect the site from falling stones. Where extra fill is placed behind an anchored retaining wall, (e.g. quay and harbour walls), the anchors are oriented horizontally and pass through the fill to connect with anchoring plates. The latter are of sufficient area to create the necessary resistance to anchor extraction, yet allowing an admissible displacement of the plate under pressure (Fig, 23-8 and 23-9). 23.2.2

Cantilever retaining walls

If a cantilever wall is anchored, the projection of the cantilever can be reduced; the necessary vertical forces can be created without the large projecting area that is otherwise required when the forces are created entirely by the weight of the fill (Fig. 23-10). Anchoring thus makes it possible to reduce the distance between the centre of gravity and the free foot of the structure. The structure may even be designed without rear cantilevers and the load centre of the anchoring forces is then placed either close to the free foot of the retaining wall, or the anchorage may pass through the wall (Fig. 23-11). This arrangement is economical because it reduces the amount of earth-moving work needed. An anchorage which passes through the wall also balances a part of the stress caused by the bending moment. The

451 Fig. 23-8. Design for an embankment wall anchored into the back-fill 1 — buttress, 2— arched retaining screen, 3 — tie bar, 4 — anchoring plate a) — vertical section, b) — plan

^\\ψΕ^Μ^^^^

Fig. 23-9. Widening of a quay by means of cantilevers built out from a wall 1 — anchors, 2 — prestress units, 3 — heads of prestressing units, 4 — retaining walls

452 Fig. 23-10. Cantilever retaining wall anchored into bedrock (Rio de Janeiro). The rear of the wall is provided with a drainage layer

11 |0*y/4v//

Fig. 23-11. Retaining wall with an interior foundation cantilever

economy is further increased where the cantilever at the interior side of the wall also forms a foundation platform for other purposes, such as roadways, etc. An interesting example of such an arrangement is the retaining wall around the Guillemine railway station area near Liege. The wall is founded on a grid mounted on piles, and the anchor tendons pass through those piles which are directly under the wall part of the structure (Fig. 23-12).

Fig. 23-12. Retaining wall for the Guillemine railway station 1 — pile grid, 2 — anchorage

453

23.2.3 Retaining walls on steep slopes Anchored retaining walls in cuttings are often constructed stage-by-stage in horizontal belts from the top downwards. Such a procedure can be followed where the soil is relatively dry and cohesive, and therefore temporarily stable up to a certain height without the need for any support. The excavation deepened to this extent is then secured with a screen of porous concrete which is supported with anchors placed into the load-bearing reinforced concrete buttresses. The anchors are oriented either horizontally or at a slight inclination, according to the degree of friction between the soil and the rear of the screen. The forces created by the prestressing of the

Fig. 23-13. Replacing an old retaining wall with an anchored structure (Brunnen, Switzerland) a) — general view of the reconstruction work, b) — drilling of boreholes from a scaffolding, c) — anchored load-bearing buttresses

454 anchors also support the weight of the partly completed supporting structure, until the entire wall has been completed. This method was used in Switzerland for the reconstruction of an old retaining wall when a highway was widened (Fig. 23-13). The old retaining wall, 42 m long, and made of rubble stone, was replaced in stages by a revetment wall anchored from the top downwards and shifted progressively deeper into the slope. The load-bearing elements were anchored into reinforced concrete buttresses which were lengthened in stages (Fig. 23-14). VSL anchoring cables 14 to 22 m long, prestressed to 650 kN were passed through a thick morainic overlayer containing large blocks, and fixed into a fractured limestone bedrock. A concrete layer made with coarse aggregate was built at the rear of the revetment wall to prevent any accumulation of water which might percolate through the slope and thus increase the pressure on the supporting structure. Again in Switzerland, a single track railway was widened to make a double track line through a long, 7 m-high cutting, without causing any interruption

Fig. 23-14. Detailed view of work on the anchored retaining wall (shown in Fig. 23-13) progressing from the top downwards (Brunnen, Switzerland) J

455

to the traffic. 122 anchors 7 to 10 m long, prestressed to 320 kN, were fixed into the non-cohesive soil of the slope. The construction of a light retaining wall with load-bearing anchored buttresses is shown in Fig. 23-15. Here, a slope had to be cut in blocky limestones for the foundation's of a new hospital. The buttresses were secured with a total of 111 VSL cable anchors from 16 to 2 0 m long, prestressed to 1.23 MN.

Fig. 23-15. Anchored revetment wall of the excavation for a hospital in Monaco

The approach cutting to a tunnel of the "Ricard Sud" motorway in France was also protected by a wall anchored into the reck. Behind the retaining wall was the tunnel of the other, northern branch of the motorway, and the anchors had therefore to be directed away from this tunnel (Fig. 23-16). The placement of anchors in the vicinity of the existing tunnel

456 Fig. 23-16. Stabilization by an anchored wall of the entrance cutting for the Richard Sud tunnel (France) 1 — tunnel for the northern motorway branch, 2 — tunnel for the southern motorway branch, 3 — retaining wall, 4,5 — anchors

resulted in the bottom of the tunnel starting to rise, as soon as test anchors were installed; this was followed by damage to the lining and the penetration of grout into the tunnel. When the circumstances of these problems were investigated, it appeared that there had been a change in the hydrogeological conditions as a result of the flushing water permeating into the rock mass, and above all, as a result of the high grouting pressure (2 MPa) applied in the fixing of the anchors. It was recommended, therefore, that water flushing be abandoned during drilling, and that the grouting pressure be reduced to 0.5 MPa. Following the completion of grouting, drainage was provided to reduce the hydrostatic pressure of the ground water [57]. This example demonstrates the correctness of the recommendations made in Chapters 14, 15 and 17, namely that when deciding on which drilling system to use, and more particularly on the optimum grouting pressure to apply, it is necessary to have a thorough knowledge of the characteristics of the ground in which the anchors are to be installed. Only then is it possible to avoid any unfavourable effects of the anchoring technique on the hydrogeological conditions within the ground, and its overall stability. The diaphragm retaining walls for the new Neasden Lane underpass on the North Circular Road in London are permanently tied back with a total of 580 Fondedile Multibell anchors (Fig. 23-17). These anchors have working loads ranging from 100 to 500 kN, and are fixed (with a safety factor of 3) in clay.

457

*td.75m

Fig. 23-17. Retaining wall of an underpass in London tied back with Fondedile Multibell anchors 1 — anchor tendon, 2 — multibells root, 3 — anchor borehole, 4 — anchor head, 5 — face of the retaining wall, 6 — retaining wall, 7 — top of the retaining wall

3d

£° 33^0

Landslides along highway cuttings in residual and colluvial soils have been a recurring problem in some areas of Brasil. The mass of the slope soil, rainfall, a slip plane in an undrained water layer collecting on the underlying rock, and the force of gravity, all occurring in various combinations, give rise to landslides of greater or lesser importance. An effective solution to this problem was found by reshaping the slope into a sequence of terraces, starting from the top, and building concrete walls against the exposed faces of each terrace with post-tensioned tie-backs to anchor the walls into the underlying rock. A 15m-high cutting near Agra dos Reis for an access road to the construction site of a nuclear power station, was stabilized by creating four terraces. These were supported by precast concrete elements measuring 1.25 x 1.00 m, each anchored with 20 to 28 m long cables (working load, 280 kN) comprising 12 wires of 8 mm diameter (Fig. 23-18). The anchors were installed in 100 mm-diameter boreholes dipping 15° from the horizontal. The holes were extended about 15 m into the rock and the fixing length was 10 m. The work was undertaken by Tecnosolo SA of Rio de Janeiro [22]. The anchored wall of the basement hall at the reconstructed central market Forum des Halles de Paris facing the Pierre — Lescot street has a special foundation on piles. On these piles of 80 cm diameter rest the pillars of the

458 wall, which is 5.30 m wide and 1.50 m thick. Between the pillars are blocking screens of 2.70 x 0.60 m cross-section to take the horizontal earth pressure [92] (Fig. 23-19).

Fig. 23-18. The upper ledge of a slope cut in landslide terrain in Brasil, secured by an anchored element wall

Fig. 23-19. Anchored wall of the basement rooms for the reconstructed central market (Forum des Halles de Paris) 1 — wall on piles, diameter 80 cm, 2 — piles, 3 — cantilever for way, 4 — anchors, a — aluvium; b — marl; c — limestone

Chapter 24 A N C H O R I N G OF C O N C R E T E

DAMS

In the construction of concrete dams there is much opportunity for adapting the structure to local conditions, so as to achieve the maximum economy of construction. The local conditions which most determine what the structure of a concrete dam will be, are the shape of the valley and the mechanical properties of the rock strata. In narrow valleys, arch dams are constructed which exploit the rock material to the optimum, making use of their static effect. Anchoring for this type of dam therefore arises only when the lateral slopes are in need of consolidatation (see Section 21.3), or when the morphological conditions are less favourable, making it necessary the arch to rest on gravitational concrete blocks. These blocks must then be secured by anchorage to the rock of the valley flanks [52]. For more open valleys, gravity dams are designed, whilst in valleys with wide flat bottoms multiple arch dams of various types are constructed. In both these types, anchoring into the bedrock substantially reduces the necessary weight of the dam, and therefore lowers the construction cost. On the 22 m-high gravity Allt-na-Lairige Dam in Scotland (Fig. 24-1), the volume of concrete required was reduced by 50 per cent, by anchoring, thus bringing down the construction cost by 17 per cent. [4], [44]. In the construction of the multiple arch dam at St. Michel in France (Fig. 24-2), where anchoring was first used in the design of a new dam, 340 m 3 of concrete were saved for every 1,000 kg of anchorage (steel), reducing the total construction cost by approximately 20 per cent.

24.1 A N C H O R I N G OF CONCRETE D A M S BY NON-PRESTRESSED A N C H O R A G E Dams should, in principle, be anchored in bedrock with prestressed anchors, because this guarantees efficient co-operation between the shear resistance of the rock and the anchoring forces from the very start of the loading of the dam. When non-prestressed anchorage is used, a partial shear failure of the dam may take place before the anchorage becomes activated. This leads to a reduction of the absolute shear resistance value of the rock, because the co-operative effect of the rock cohesion, c, will be lost and the static coefficient of static friction, / , will be replaced by the coefficient of

460 Fig. 24-1. Allt-na-Lairige Dam (Scotland) 1 — bar bundles of the anchors, 2 — anchoring shaft, 3 — anchor root, 4 — anchor head, 5 — couplings

Α-Α' A

Γ

?

Π

c3

(

\\ \v

SJQ

Fig. 24-2. St. Michel multiple dam (France) 1 — cables embracing the buttresses, 2 — saddle heads

461

kinetic friction, fr, which has a lower value (see also Section 21.1). For these reasons, non-prestressed bar anchors are only used for dams under a height of 6.5 m (according to J. K. Wilkins and J. Fidler [223]). Only in exceptional cases is non-prestressed anchorage used in the construction of taller dams. The 54 m-high Aventino Dam in Italy (in the Apennines) was secured with non-prestressed steel bars 24 to 30 mm in diameter. These were placed in 95 mm-diameter boreholes 10 m deep, and grouted along their entire lengths [5]. The upper ends of the anchoring bars projected 2.0 m into the dam buttresses, which at this level were strongly reinforced with 24 mm-diameter steel bars (Fig. 24-3). The bedrock of the A-A'

^

B-B' B'

Fig. 24-3. Anchoring of the Aventino Dam (Italy)

dam consisted of low permeability beds of calcareous breccias, more or less cemented breccias of calcareous marlstones, and clay beds with interlayers of limestone and sandstone. In order to secure the dam, 1,335 boreholes had to be drilled, with a total length of 12,149 m. The total amounts of steel and cement used were 78,000 kg, and 352,000 kg, respectively. The reinforced parts of the buttresses at the footings had an overall volume of 12,000 m 3 and consumed 486,000 kg of steel. Although the measures adopted have proved reliable, it is clear from the volumes of materials used that this success was bought at great cost. The same results could have been obtained with rather less material and effort, had prestressed anchorage been used for fastening the buttresses to the bedrock. 24.2 A N C H O R I N G OF CONCRETE D A M S BY PRESTRESSED ANCHORAGE In the calculation of the anchoring forces required, the stability of the dam against horizontal displacement at the foot is considered first. The stability of the dam against overturning, or against shear failure along the

462

critical subsoil surface, is assessed in the second stage, and if neccessary, the computed value of the anchoring forces is modified to guarantee the stability of the dam against any possible failure. 24.2.7

Design of anchorage for gravity dams

The anchoring forces required to secure a concrete dam against horizontal displacement are determined as follows (Fig. 24-4):

Fig. 24-4. Forces considered in the anchoring of gravity dams against horizontal displacement

The safety factor for horizontal displacement of the dam is determined by the relation:

1

i

m, = -*—

7b -

ih1 2

h

+ Kjtgcp =

[{n-T)+lf\^

where ms = safety factor for horizontal displacement of the dam, i = — = slenderness ratio, h s = width of the dam base [m], h = height of the static triangle of the dam [m], yb = density of concrete [kN/m 3 ], u = uplift pressure (the value of u depends on the quality of the bedrock and the function of the grout curtain) [kN], K = weight of the concrete crest [kN]. If it is found that the calculated safety factor, w s , is too small, it is increased by introducing anchoring forces, P, as follows:

463

m,

-['(»- τ) +