Amund Bruland Hard Rock Tunnel Boring ...

Doctoral theses at NTNU 1998:81

Amund Bruland Hard Rock Tunnel Boring Vol. 6 of 10

Performance Data and Back-mapping

NTNU Trondheim Norwegian University of Science and Technology Doctoral thesis for the degree of doktor ingeniør Faculty of Engineering Science and Technology Department of Civil and Transport Engineering

0

1

2

3

4

PREFACE

1

GENERAL

3

0.1 Project Reports about Hard Rock Tunnel Boring

3

0.2 Field Performance Studies

6

MACHINE PERFORMANCE

8

1.0 Introduction

8

1.1 Net Penetration Rate

9

1.2 Gross Thrust

12

1.3 Torque

14

1.4 Machine Utilisation

15

1.5 Weekly Advance Rate

20

1.6 Parameter Summary

22

CUTTER CONSUMPTION

23

2.0 Introduction

23

2.1 Consumption per Position

24

2.2 Instantaneous Consumption

28

2.3 Ring Wear

34

PENETRATION TESTS

40

3.1 Test Procedure

40

3.2 Penetration Curve

44

3.3 Cutter Coefficient

47

3.4 Specific Energy

50

CHIP ANALYSIS

51

4.1 Test Procedure

51

4.2 Chip Size

54

4.3 Sieve Curve

58

4.4 Kerf Depth Factor

59

4.5 Crack Growth

61

4.6 Rock Properties

62

5

BACK-MAPPING

64

5.0 Introduction

64

5.1 Mapping Procedures

65

5.2 Aggregation of Mapping Data

71

5.3 Rock Sampling

77

5.4 Achieved vs. Predicted Performance

78

APPENDICES

79

A.

Previous Editions

79

B.

Research Partners

80

C. List of Parameters

81

D.1 Mapping Sheet ±135°

86

D.2 Mapping Sheet ±180°

87

D.3 Shift Log

88

D.4 Cutter Change Log

89

D.5 Sieve Curve Sheet

90

E.

91

Penetration Test Spreadsheet

PREFACE

HARD ROCK TUNNEL BORING Performance Data and Back-mapping Project Report 1E-98 The report is one of six reports about hard rock tunnel boring: • • • • •

1A-98 HARD ROCK TUNNEL BORING Design and Construction 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear 1C-98 HARD ROCK TUNNEL BORING Costs 1D-98 HARD ROCK TUNNEL BORING Geology and Preinvestigations 1E-98 HARD ROCK TUNNEL BORING Performance Data and Back-mapping • 1F-98 HARD ROCK TUNNEL BORING The Boring Process In addition, HARD ROCK TUNNEL BORING Background and Discussion gives general information about the basis of the above listed reports. Combined with the other reports in the Project Report Series from the Department of Building and Construction Engineering at NTNU, the reports present an updated and systematised material on rock excavation and tunnelling to be used for: • • • • •

Economic dimensioning Choice of alternative Time planning Cost estimates, tender, budgeting and cost control Choice of excavation method and equipment.

A list of available Project Reports may be requested from the Department of Building and Construction Engineering at NTNU. The advance rate, cutter wear and excavation cost models also exist as a WINDOWS programme. The report is prepared by Amund Bruland and is part of his dr.ing thesis about hard rock tunnel boring.

1

PREFACE

The reports listed above describes a comprehensive model developed at NTNU The model covers the complete tunnel boring process from the early planning stage through preinvestigations, time and cost estimates, tunnel excavation and finally acquisition and treatment of experience data. The models and data presented in the reports are meant to be a practical tool for owners, consultants and contractors, more than a theoretical analysis of the tunnel boring process. The project has been granted financial support by our external research partners, see list in Appendix. For reference, registration and similar, we ask for the following: NTNU-Anleggsdrift (1998): Project Report 1E-98 HARD ROCK TUNNEL BORING Performance Data and Back-mapping. When copying from the report, the source should be stated.

Trondheim, September 2000

Odd Johannessen Professor

Contact address: Amund Bruland Department of Building and Construction Engineering, NTNU N-7491 Trondheim NORWAY Telephone +47 73 59 47 37 Fax +47 73 59 70 21 e-mail [email protected] Internet http://www.bygg.ntnu.no/batek/batek.htm

2

0. GENERAL 0.1


PROJECT REPORTS ABOUT HARD ROCK TUNNEL BORING 1E-98 The report provides methods and data to be used during follow-up and analysis of TBM performance and for back-mapping of engineering geological properties of bored tunnels.

The report treats the following items: • Chapter 1 describes a system for recording, averaging and normalisation of machine performance data such as net penetration rate, gross advance rate, machine utilisation, etc. • In Chapter 2, methods for recording and analysis of the cutter wear of individual cutters and the cutterhead as a whole are presented. • Chapter 3 describes how to perform and analyse penetration tests. • Various measurement procedures and how to analyse the rock breaking process using the largest chips are treated in Chapter 4. • Procedures for engineering geological back-mapping and normalisation of mapping data are described in Chapter 5. Project Report 1E-98 is partly based on the project reports 1-76, 1-79, 1-83, 1-88 and 1-94, all published by the Department of Building and Construction Engineering at NTNU. The report presents updated and revised information from the previous reports as well as experience from recent tunnelling projects. Appendix A shows a list of previous editions of the HARD ROCK TUNNEL BORING report.

Other Reports The Project Report 1A-98 HARD ROCK TUNNEL BORING Design and Construction describes general design parameters such as tunnel profile, tunnel inclination and curve radius. Some features of various tunnel types like water, sewage, road

3

0. GENERAL


and rail tunnels are treated. Transport, ventilation and other necessary service systems are presented. The Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear provides methods and necessary data for estimation of time consumption and cutter wear for tunnel boring. Geological parameters and machine factors of significance for the penetration rate and the cutter wear are presented briefly. The Project Report 1C-98 HARD ROCK TUNNEL BORING Costs presents models and data for estimation of tunnel excavation costs and total construction costs. The Project Report 1D-98 HARD ROCK TUNNEL BORING Geology and Site Investigations treats the rock mass parameters of the model in more detail. Site investigations and building of an engineering geological model adapted to the estimation models for penetration rate and excavation costs are treated closely. Project Report 1F-98 HARD ROCK TUNNEL BORING The Boring Process covers rock breaking and chipping, machine factors affecting performance, boring in fractured rock mass, and various topics of cutter wear.

Use of the Estimation Models The estimation models are aimed at being used through several stages in a project: • • • • •

Preliminary and feasibility studies Project design and optimisation Tendering and contract Construction Possible claims.

The estimation models for Hard Rock Tunnel Boring should be used with care. Combined with other estimation models in the Project Report Series from the Department of Building and Construction Engineering, the Hard Rock Tunnel Boring reports provide a reliable and practical tool to be used for:

4

0. GENERAL

• • • • • •


Estimating net penetration rate and cutter wear Estimating time consumption and excavation costs, included risk Assess risk with regard to variation in rock mass boreability or machine parameters Establish and manage price regulation in contracts Verify machine performance Verify variation in geological conditions.

Background The estimation models are based on job site studies and statistics from tunnelling in Norway and abroad, including more than 35 job sites and more than 250 km of tunnel. The data have been systematised and normalised. The results are regarded as being representative for well organised tunnelling. A more detailed treatment of the background and the basis for the Hard Rock Tunnel Boring estimation models is found in HARD ROCK TUNNEL BORING Background and Discussion.

5

0. GENERAL


0.2 FIELD PERFORMANCE STUDIES The performance data of the machine are combined with the cutter wear data and data from the engineering geological back-mapping to be able to evaluate the performance of the tunnel excavation with regard to the geological conditions along the tunnel. Field performance studies at a TBM tunnel aim to provide information and knowledge to be used for several purposes: • To improve the tunnel boring operations at the site. • To verify the tunnelling performance against the models and assumptions used in the planning process. • To increase the knowledge and expertise of the tunnelling organisation. • To provide data for further development of the TBM and tunnelling technology in general. • To provide data for development and improvement of the general prediction models related to TBM tunnelling, to be used by planners, contractors, manufacturers, etc. The field performance studies should build a database that describes the geology and the tunnelling performance by numbers. Each record of the database should represent tunnel sections with approximately constant basic parameters, e.g. rock type, tunnel direction, cutter thrust, cutterhead RPM. An example of sectioning of the database is shown in Figure 0.1. Z o n e 1

Z o n e 2

R a n g e 1

R a n g e 2

R a n g e 1

R a n g e 2

R a n g e 3 R a n g e 3

R a n g e 1 R e c o rd 1

Z o n e 3

G e o lo g y

R a n g e 4 R a n g e 4

R a n g e 5

2

R a n g e 2 3

4

T B M p a ra m e te rs P e n e tr a tio n ra te

5

C u tte r c o n s u m p tio n 6

D a ta b a s e

Figure 0.1 Sectioning of the back-mapping and follow-up database. 6

0. GENERAL


In this process, the geology and the machine performance are related to the chainage along the tunnel. Hence, an accurate and close marking of the chainage is recommended, e.g. with minor indicators at each metre and major indicators at each 10 metres. Furthermore, the performance data and the geological back-mapping are normalised and included in the general database of hard rock tunnel boring to form the basis for improving and extending the prediction models for time and cost estimates of TBM tunnelling. The data collection and the development of the prediction models is a more or less continuous process since the models are purely empirical and new input in the form of e.g. machine design, cutter material and geological conditions is constantly available. The field performance studies and data collection described in this report are consistent with the basic follow-up program performed by the Department of Building and Construction Engineering at NTNU at various tunnel boring sites. At a given tunnel project, it will often be of interest, or even necessary, to perform additional or more detailed observations and analyses of selected processes, equipment, parameters or similar. When planning and performing the follow-up of a specific tunnelling project, one should keep in mind that detailed studies will get an increased value if combined with general follow-up as described in Sections 1.0, 2.0 and 5.0 of this report.

7

1. MACHINE PERFORMANCE

1.0 Introduction

1.0 INTRODUCTION The machine performance is continuously recorded through the shift logs and/or log files from the onboard computer. Specific tests such as penetration tests give valuable and more detailed information about the machine performance. The logs should as a minimum provide data to be able to calculate the instantaneous and average of: • • • • •

Net penetration rate (mm/rev and m/h) Applied cutter thrust (kN/cutter) Torque (kNm) and cutter coefficient Machine utilisation (%) Gross advance rate (m/week).

An example of a shift log is given in Appendix D.

8



1.1 NET PENETRATION RATE The average net penetration rate is calculated based on systematic and continuous recording of the chainage and the machine hours, recorded e.g. at the end of each shift. For the accuracy of the records it is important that the machine hour meter measures boring hours. Hence, it must only be activated by a combination of cutterhead rotation and cutterhead thrust. It is convenient to base the follow-up work on the average net penetration rate over one shift, see Figure 1.1. When an onboard computer is used for data acquisition and logging, shorter intervals like the stroke length may be used when the average net penetration rate is calculated. The penetration rate calculated from the logging system (shift log or computer) should be verified by penetration rate measurements, e.g. over one stroke length. 5 m/h 4

3

2

1

0 2450

2460

2470

2480

2490

2500 Chainage, m

Figure 1.1 Net penetration rate at the Meraaker Project, calculated from shift logs. dtbm = 3.5 m. Courtesy of NCC Eeg-Henriksen Anlegg AS and Veidekke ASA.

The detailed net penetration records are compared to the engineering geological backmapping at a detailed level. Figure 1.1 may be explained as follows: 9



• The relative high net penetration rate from chainage 2450 to 2460 and from chainage 2490 to 2500 may have various reasons, of which the most likely are: ♦ Higher average cutter thrust ♦ Higher degree of fracturing ♦ Influenced by a Marked Single Joint ♦ Weaker rock type The average penetration rate must be calculated using the total machine hours used to bore the actual tunnel length, and not as an arithmetical average of the penetration rates of the subsections (e.g. shifts) constituting the actual tunnel length.

Im =

∑l ∑T

j

bj

Im Inj lj Tbj

=

∑l

j

∑ (l j / I nj )

(m/h)

[1.1]

= average net penetration rate (m/h) = net penetration rate for subsection j (m/h) = length of subsection j (m) = net time (machine hours) used to bore subsection j (h)

It is convenient to average the net penetration rate over time periods of equal length (e.g. weeks, see Figure 1.2) or over tunnel sections, each section with more or less constant geological conditions.

10



9 m/h 8 Cumulated average 7 6 5 4 3 2 1 0 38

42

46

50

2

6

10

14

18

22

26

30

34

Week no. of 1991/1992

Figure 1.2 Average net penetration rate at the Meraaker Project, calculated from shift logs. dtbm = 3.5 m, Lt = 10 km. Courtesy of NCC Eeg-Henriksen Anlegg AS and Veidekke ASA.

11


1.2 Gross Thrust

1.2 GROSS THRUST The prediction model in the Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear uses the gross average thrust per cutter as a measure for the applied cutter force. Gross average thrust means: • The total gross thrust is the gross thrust force from the thrust or propel cylinders, including friction or drag of the cutterhead and the cutterhead support system and other losses of thrust. • The total gross thrust is divided by the number of cutters on the cutterhead to get the gross average thrust per cutter. The above system is meant to simplify and standardise the data acquisition and treatment. Individual and net cutter forces are very difficult to measure and will introduce additional uncertainty in the performance data analyses and in the prediction model. It may however be necessary to instrument individual cutters to measure the cutter forces for research and development purposes. The applied gross thrust is calculated from the applied hydraulic pressure in the propel cylinders, as shown in Section 3.2, Equation [3.5]. The applied hydraulic pressure is recorded continuously by the onboard computer or at regular intervals in the shift log by readings of the propel pressure gauge. The shift log should record the propel pressure at least two times during one stroke length. The applied thrust will oscillate somewhat. Since the prediction model for net penetration rate is based on the applied gross average cutter thrust, it is important to use the average reading of the pressure gauge and not the maximum deflection of the pointer. When the applied thrust is averaged over a given tunnel length, the instantaneous thrust is weighted by time.

M Bm =

∑ (M ⋅ T ∑T Bj

bj

)

(kN/cutter)

[1.2]

bj

12


1.2 Gross Thrust

MBm = gross average thrust over a given tunnel length (kN/cutter) MBj = gross average thrust over subsection j (kN/cutter) Tbj = net time in machine hours for boring subsection j (h)

Tbj =

lj Inj

lj I nj

(h)

[1.3]

= length of subsection j (m) = net penetration rate of subsection j (m/h)

The gross thrust should be averaged over the same time periods and tunnel sections as the net penetration rate.

13


1.3 Torque

1.3 TORQUE The applied torque TB is calculated from the applied amperage of the cutterhead drive motors. Equation [1.4] is applicable for 3-phase motors.

TB =

U B ⋅ I B ⋅ 3 ⋅ (cos φ ⋅ η ) ⋅ 60 ⋅ nm 2 ⋅ π ⋅ 1000 ⋅ RPM

(kNm)

[1.4]

UB = applied operating voltage of the cutterhead drive motors (V) = applied amperage (A) IB cosφ⋅η = efficiency of the motors at the given amperage, see Section 3.3 (if unknown, cosφ⋅η = 0.8 is a good approximation) = number of motors on the cutterhead nm RPM = cutterhead revolutions (rev/min) The applied amperage is recorded continuously by the onboard computer or at regular intervals in the shift log by readings of the ampere meters. The shift log should record the amperage at least two times during one stroke length. The applied amperage will oscillate somewhat. Since the prediction model for necessary torque is based on the average amperage, it is important to use the average reading of the ampere meter and not the maximum deflection of the pointer. When the applied amperage is averaged over a given tunnel length, the instantaneous amperage is weighted by time.

I Bm =

∑ (I ⋅ T ∑T Bj

bj

)

(A)

[1.5]

bj

IBm IBj Tbj

= average amperage over a given tunnel length (A) = amperage over subsection j (A) = net time for boring subsection j (h), see also [1.3]

The amperage should be averaged over the same time periods and tunnel sections as the net penetration rate. 14



1.4 MACHINE UTILISATION The machine utilisation is defined as the boring time in percent of the total available shift or working hours. The machine utilisation is calculated for given time periods (e.g. week, month) or for given tunnel sections (e.g. geological zones, total tunnel length). The calculation of the machine utilisation is entirely based on data from the shift logs recorded by the TBM operator. In some cases, it is difficult to state the appropriate cause of stop in the boring. In the future, use of the onboard computer for semiautomatic data recording and treatment will improve the reliability and objectivity of the shift log data. A digital shift log will also improve the possibilities and the results of back-analyses of the machine performance. The shift log records: • Chainage and machine hours. • Basic machine operation parameters (applied thrust, torque and cutterhead RPM, if variable). • Time consumption of main activities. • Main causes of stops in the boring. An example of a well-suited shift log for hard rock tunnel boring is shown in Appendix D. The grouping of the activities may vary, but the following is recommended to be included in the main activities: • • • • • • • • • • •

Boring, including regripping Tbr Cutter change and inspection Tc Repair and service of the TBM Ttbm Repair and service of the backup equipment Tbak Muck transport (if separate transport system such as continuous conveyor) Tmt Other transport (personnel, equipment, materials) Tot Supply of water, electricity and ventilation Tw Surveying Ts Personal time such as travel, change of crews, lunch breaks, etc. Tp Rock support, water inflows and other geological stop causes Trs Other activities Toa

15



The main activities may be subdivided into even more specific activities such as for rock support in a continuously lined tunnel in weak and water-bearing rock mass. ♦

Rock support • Probe drilling • Pregrouting • Segment erection • Back-filling of the segmental lining • Water inflow • Other rock support.

One must however remember that the shift log is recorded by the TBM operator who normally gives priority to the tunnel excavation operations. Hence, to ensure the quality of the recorded data, the shift log must not be too complex asking detailed information of little relevance. The causes of stops in the boring may be entered as comments if necessary. It is not convenient to record boring and regripping separately in the shift log. When calculating the machine utilisation, one has to combine the recorded machine hours (net boring time) and the recorded time consumption of boring, including regripping as shown in [1.6] and [1.7]. The equations are based on a time period of one week. Tb = t 2 − t1

Tb t1 t2

[1.6]

= net time for boring (h/week) = machine hours at the start of the actual week (h) = machine hours at the end of the actual week (h)

Tr = Tbr − Tb

Tr Tbr

(h /week )

(h /week )

[1.7]

= time for regripping (h/week) = time for boring, including regripping, per week, from the shift log (h/week) 16



To be able to compare the achieved machine utilisation with the normalised model in Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear, the time consumption for rock support work is excluded from the total working hours to make up the available excavation time. The reason for this is that the type and amount of rock support work is mainly dependent on the site geology and much less by the tunnel boring system. Tex = Tsh − Trs

Tex Tsh Trs

ua =

ua Tb Iu In

[1.8]

(h /week )

= available time for tunnel excavation (h/week) = shift hours (h/week) = time for rock support work (h/week) Tb I u ⋅ 100 = Tex I n ⋅ Tex

[1.9]

(%)

= achieved machine utilisation per week (%) = time for boring (h/week) = weekly advance rate (m/week) = average net penetration rate per week (m/h)

The utilisation must also be corrected for weekly working hours (shift hours) other than 100 h/week.

un = ua ⋅

un Teh Tsh

1

Tsh Teh

(%)

[1.10]

= normalised machine utilisation per week of 100 shift hours (%) = effective working hours (h/week)1 = shift hours (h/week)

Figure 4.2 in Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter

Wear.

17



An example of a plot of the weekly machine utilisation and cumulated machine utilisation is shown in Figure 1.3. 70 % 60

Cumulated average

50

40

30

20

10

0 38

42

46

50

2

6

10

14

18

22

26

30

34

Week no. of 1991/1992

Figure 1.3 Machine utilisation at the Meraaker Project. dtbm = 3.5 m, Lt = 10 km. Courtesy of NCC Eeg-Henriksen Anlegg AS and Veidekke ASA. The cumulated average machine utilisation for a given tunnel length or period of time is calculated directly from the cumulated time consumption for the given period as in [1.11]. If the weekly working hours varies from week to week, the cumulated utilisation may be calculated by [1.12], where the utilisation is weighted by working hours per week. Non-productive stops like vacations are not included in the cumulated average.

um = ua =

∑T ∑T

b

⋅ 100

(%)

[1.11]

ex

um =

∑ (u a ⋅ Tsh ) ∑ Tsh

(%)

[1.12]

18



The total use of time distributed on the various tunnelling operations may be illustrated by pie charts as the example in Figure 1.4.

Other 4.6 Water etc. 10.0

Transport 9.7

Boring 40.2

Backup 4.1

TBM 5.4

Cutters 15.3 Regripping 10.7

Figure 1.4 Tunnelling time consumption in per cent distributed on various operations at the Meraaker Project. dtbm = 3.5 m, Lt = 10 km, Im = 6.4 m/h. Courtesy of NCC Eeg-Henriksen Anlegg AS and Veidekke ASA.

19



1.5 WEEKLY ADVANCE RATE

m/week

450 400 350

Cumulated average

300 250 200 150 100 50 0 38

42

46

50

2

6

10

14

18

22

26

30

34

Week no. of 1991/1992

Figure 1.5 Weekly advance rate at the Meraaker Project. dtbm = 3.5 m, Lt = 10 km. Courtesy of NCC Eeg-Henriksen Anlegg AS and Veidekke ASA.

The cumulated average weekly advance for a given tunnel length or period of time is calculated directly from the bored tunnel length and the number of productive weeks in the period. The cumulated average may also be calculated by the cumulated averages as in [1.14].

Iu =

l1 l2 nu

l 2 − l1 nu

(%)

[1.13]

= chainage at the start of the period (m) = chainage at the end of the period (m) = number of productive weeks of the period

20


I u = u m ⋅ Tex ⋅ I m

(m/week)


[1.14]

To be able to compare the achieved weekly advance rate with the normalised model in Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear, the advance rate must be related to 100 working hours available for tunnel excavation per week, and also be corrected for weekly working hours (shift hours) other than 100 h/week.

I un = u m ⋅ 100 ⋅ I m ⋅

Tsh Teh

(m/week)

[1.15]

21


1.6 Parameter Summary

1.6 PARAMETER SUMMARY It is convenient to aggregate the detailed information from the shift logs into a more readable form. A visual presentation as shown in Figure 1.6 gives a quick and very informative overview of the operation of the machine, as well as providing a basis for analyses of the machine performance with regard to the engineering geological backmapping.

N e t p e n e tr a tio n r a te , m /h

The aggregated performance data are also a good check for large errors in the backmapping. It is however not recommended to bring the summary into the tunnel while performing the back-mapping. This may bias the classification of the rock mass degree of fracturing or other parameters of the mapping.

5 4 3 2 1 0

T h ru s t, k N /c u tte r

2 5 0

2 0 0

1 5 0

T o rq u e , A

2 5 0

2 0 0

1 5 0 2 4 5 0

2 4 6 0

2 4 7 0

2 4 8 0

2 4 9 0 2 5 0 0 C h a in a g e , m

Figure 1.6 Summary of averaged machine performance parameters over shifts, based on data from the shift logs. 22

2. CUTTER CONSUMPTION

2.0 Introduction

2.0 INTRODUCTION The cutter consumption is recorded through the cutter change and inspection log in the tunnel and the cutter repair log at the cutter shop. Specific tests such as the wear progress of specific cutter rings and positions, steel quality measurements, etc., give valuable and more detailed information about the cutter wear process. The logs should as a minimum provide data to be able to compute: • • • • •

Instantaneous and average cutter wear for the cutterhead (h/c, m/c and sm3/c) Cutter consumption for each cutter position (e.g. no. of rings) Reason for change (e.g. ring wear, oil leakage, blocked bearings) Type and extent of the wear of each ring Consumption of spare parts.

To be able to trace the individual cutter and record the necessary data for the abovementioned calculations, each cutter must have a unique identity, e.g. a number. The cutter change log must be related to the tunnel chainage and the machine hours. An example of a cutter change log is given in Appendix D. If swapping of partly used cutters between positions (e.g. from outer to inner gauge) is done, this must be noted in the cutter change log and corrected for in the calculations of the cutter consumption. To be able to calculate the instantaneous consumption towards the end of the tunnel, one should record the wear state of all cutters when the boring is finished. The state is recorded as wear height or as a rough estimate of remaining ring life in per cent. If the boring of the tunnel starts with partly used cutter rings, the wear state of the cutters should be recorded at the start of the boring in the same way.

23



2.1 CONSUMPTION PER POSITION The cutter consumption per position is used to: • Evaluate the cutterhead design, e.g. with regard to placement of the individual cutter. • Calculate the cutterhead factor fD and the correction factor for TBM diameter with regard to cutter consumption, see Figure 3.2 of Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear.

No. of rings used

35 Other Chipped ring Oil leakage Blocked Wear

30

25

20

15

10

5

0 1

2

3

4

5

6

7

8

9

10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 Cutter position no.

Figure 2.1 No. of rings used per cutter position at the Meraaker Project. dtbm = 3.5 m, Lt = 10 km, Ntbm = 25. Courtesy of NCC Eeg-Henriksen Anlegg AS and Veidekke ASA. Figure 2.1 shows two important features of the given cutterhead design: • The positions 12 through 17 are exposed to the highest loads, indicated by the number of cutters replaced for reasons other than wear. The load of the cutters may be reduced by adding one cutter in this area (if possible).

24



• The centre cutter positions no. 2 and 4 should be examined to find a better design, indicated by the number of cutters replaced for reasons other than wear. In the actual tunnel, the positions could be relieved by replacing the cutter in position 5 more often. The cutterhead factor is calculated by normalising the cutter ring consumption per position to a cutterhead with radius 1 and an average cutter consumption of 1 ring per position as shown below.

rri =

Ni N tbm

[2.1]

rri = relative radius of cutter position no. i Ni = cutter position no. i Ntbm = number of cutter positions on the cutterhead

H ri =

∑ H ni

N tbm ⋅ H ni

Hri Hni

= relative cutter ring life of position no. i = number of cutter rings used at position no. i

fD =

∑H

fD

= cutterhead factor

kD =

fD f D0

kD fD0

= correction factor for TBM diameter with regard to cutter ring life = 1.133, i.e. the cutterhead factor of the reference cutterhead (dtbm = 3.5 m, dc = 394 mm, Ntbm = 27)

ri

N tbm

[2.2]

[2.3]

[2.4]

25



Position No.

No. of Rings Used

Relative Position

Relative Life

1

11

0.04

1.8400

2

13

0.08

1.5569

3

11

0.12

1.8400

4

13

0.16

1.5569

5

12

0.20

1.6867

6

11

0.24

1.8400

7

11

0.28

1.8400

8

13

0.32

1.5569

9

14

0.36

1.4457

10

19

0.40

1.0653

11

22

0.44

0.9200

12

29

0.48

0.6979

13

33

0.52

0.6133

14

29

0.56

0.6979

15

30

0.60

0.6747

16

32

0.64

0.6325

17

28

0.68

0.7229

18

25

0.72

0.8096

19

25

0.76

0.8096

20

24

0.80

0.8433

21

22

0.84

0.9200

22

22

0.88

0.9200

23

21

0.92

0.9638

24

19

0.96

1.0653

25

17

1.00

1.1906

Total

506

Cutterhead factor Correction factor for TBM diameter

Table 2.1

28.7098

28.7098 = 25 1.1484 kD = = 1.133

fD =

1.1484 1.01

Calculation of cutterhead factor and correction factor for TBM diameter with regard to cutter consumption. Data from Figure 2.1.

26



Table 2.1 shows the calculation of the cutterhead factor fD and the correction factor kD for the data in Figure 2.1.

R e la tiv e c u tte r life

Figure 2.2 shows observed and normalised cutter replacement curves (relative cutter life) for various TBM diameters. The integral of the curves gives the cutterhead factor.

6 .0

5 .0

4 .0

3 .0

2 .0

d

1 .0

0 .2

0 .4

0 .6

0 .8

tb m

= 6 .5 m 4 .5 m 3 .5 m

1 .0 R e la tiv e c u tte r p o s itio n

Figure 2.2 Normalised cutter replacement curves for varying TBM diameter.

27



2.2 INSTANTANEOUS CONSUMPTION The instantaneous cutter consumption along the tunnel is calculated between cutter changes as illustrated in Figure 2.3. The figure shows the calculation model for cutter wear in cutter/m. The same model is used to calculate cutter/h, simply by substituting the chainage (in metres) for each cutter change with the corresponding machine hours. C h a in a g e m C h a n g e n o . N c

1 2 0

2 0 0

2 9 0

3 4 0

4 1 0

1

2

3

4

5

w

1 ,3

=

3 )

w

2 ,3

= w

1 ,3

w

1 ,2

=

2 )

w

2 ,2

= w

1 ,2

w

1 ,1

=

1 )

w

3 ,3

= w

w

4 ,3

= N /A

3 ,2

= N /A

w

4 ,2

= N /A

=

w

4 ,1

= w

1 ,3

C u tte r p o s itio n , N

i

3 w

2 w

2 ,1

=

4 )

w

3 ,1

m 3

= å w

5 )

3 ,1

1

T o ta l w e a r [c /m ] w

3

T o ta l life [m /c ] H

m 1

= å w

m 1

i= 1

1 ,i=

= w

1 m 1

0 .0 2 2 9

w

= 4 3 .6 7

1 0 0

3

H

m 2

2 0 0

= å w i= 1

2 ,i=

m 2

=

0 .0 2 1 5

3

w

3 ,i

i= 1

3

w

= å w

m 4

w

= N /A

i= 1

1 m 2

= 4 5 .5 1

H

m 3

= N /A

H

S ta r t b o r in g w

1 ,1

=

1 = 0 .0 1 2 5 c /m 2 0 0 -1 2 0

3 )

w

1 ,3

=

1 = 0 .0 0 4 5 c /m 3 4 0 -1 2 0

2 )

w

1 ,2

=

1 = 0 .0 0 5 9 c /m 2 9 0 -1 2 0

4 )

w

2 ,1

=

1 = 0 .0 1 1 1 c /m 2 9 0 -2 0 0

N /A

= N /A

4 0 0 C h a in a g e , m

3 0 0

1 )

m 4

4 ,i=

5 )

w

3 ,1

=

1 = 0 .0 0 8 3 c /m 4 1 0 -2 9 0

= c u tte r p o s itio n r e p la c e d

Figure 2.3 Calculation of cutter consumption in cutter/m for a theoretical cutterhead with three cutters.

28



It must be noted that the terms cutter wear, cutter consumption, cutter/m and cutter/h refer to the consumption of cutter rings only, and not of complete cutter assemblies. The minimum requirements to the cutter change log are that it shows the following for each cutter change: • Cutter position(s) replaced • Chainage in metres • Machine hours Preferably, the cutter change log should include information on cutter identification, reason for change, etc. See Appendix D. The following definitions are recommended to avoid confusion: Cutter change

Defined as the operation of replacing one or more worn or damaged cutters at a given chainage. Cutter change no. 1 is defined as the start of the boring.

Cutter replacement

Defined as the replacement of a worn cutter with a new or rebuilt cutter (new ring as a minimum) at a single position during a cutter change.

The basic concept used to calculate the cutter consumption based on the cutter change logs is shown in Figure 2.3. The detailed model will not be treated here, but is available as a PC program from the Department of Building and Construction Engineering at NTNU. Table 2.2 shows the minimum necessary input data of the PC program as recorded in the data file. "1" means that the position has been replaced at the actual cutter change and "0" means no cutter replacement. The data in Table 2.2 are from a tunnel bored in extremely hard and strong rock with medium abrasivity. This is reflected in the frequent cutter changes and very low cutter life shown in Tables 2.2 and 2.3 and in Figure 2.4.

29



Cutter Change No.

Chainage (m)

Machine hours (h)

Positions 1 - 25

1

122.3

2186.5

1111111111111111111111111

2

164.4

2208.4

0000000000100000000000000

3

179.5

2214.4

1111100000011111111111100

4

181.7

2215.0

1010000000000000000000000

5

195.6

2223.1

0000000000001111111100000

6

219.0

2238.0

0000000000000011111100000

7

228.0

2242.8

0000000000001001000000000

8

232.6

2244.6

0000000000000111111111100

9

244.4

2252.2

0000000000000000010110000

10

265.0

2264.1

0000000000000111111110000

11

288.0

2277.1

0000000000000000011110000

12

297.8

2283.9

1111111111111111111111111

13

305.1

2290.4

0000000000000011111000000

14

311.1

2293.0

0000000000001000000100000

15

316.5

2296.8

0000000000000011100000000

16

325.0

2300.5

0000000000000001111111000

…

…

…

…

Table 2.2

Data file for calculation of cutter consumption along the tunnel. TBM with 25 cutters.

Table 2.3 and Figure 2.4 show results of the calculation of the instantaneous cutter consumption based on the data in Table 2.2.

30

Table 2.3

179.5

181.7

195.6

219.0

228.0

232.6

244.4

265.0

288.0

297.8

305.1

311.1

316.5

325.0

164.4

179.5

181.7

195.6

219.0

228.0

232.6

244.4

265.0

288.0

297.8

305.1

311.1

316.5

(m)

(m)

164.4

Chainage

Chainage

122.3

To

From

14.9

14.9

13.7

9.4

14.3

21.9

21.7

17.4

12.3

13.9

20.5

14.8

6.2

27.9

26.7

(m /c)

3

Hf

0.76

0.84

0.81

0.75

0.96

1.32

1.32

1.11

0.60

0.75

1.30

.0.84

0.22

1.44

1.38

(h/c)

Hh

1.55

1.55

1.43

.98

1.48

2.28

2.26

1.81

1.28

1.45

2.13

1.53

.65

2.90

2.77

(m/c)

Hm

0.646

0.646

0.700

1.022

0.674

0.439

0.443

0.552

0.781

0.692

0.469

0.652

1.544

0.345

0.361

(c/m)

wm

8.5

5.4

6.0

7.3

9.8

23.0

20.6

11.8

4.6

9.0

23.4

13.9

2.2

15.1

42.1

(m)

Length

Interval

7

3

2

5

25

4

8

3

10

2

6

8

2

17

1

Changed

Cutters

28

12

8

20

100

16

32

12

40

8

24

32

8

68

4

Changed

%

2.30

1.42

2.31

1.12

1.44

1.77

1.73

1.55

2.56

1.87

1.57

1.72

3.67

2.52

1.92

(m/h)

In

2.04

1.85

1.76

1.31

1.54

1.72

1.70

1.63

2.12

1.93

1.64

1.83

2.91

2.02

2.01

(m/h)

Ic

0.89

1.30

0.76

1.17

1.07

0.97

0.98

1.05

0.83

1.03

1.05

1.06

0.79

0.80

1.05

Ic /In

2. CUTTER CONSUMPTION 2.2 Instantaneous Consumption

Instantaneous cutter consumption along the tunnel, based on the data in Table 2.2.

31



In Table 2.3, Hf, Hh, Hm and wm are calculated according to Figure 2.2. In is the achieved penetration rate calculated from the chainage and the machine hours.

I n ,i =

lc,i lc,i+1 hc,i hc,i+1

l c ,i +1 − l c ,i hc ,i +1 − hc ,i

(m/h)

[2.5]

= chainage of cutter change no. i = chainage of cutter change no. i+1 = machine hours of cutter change no. i = machine hours of cutter change no. i+1

Ic is the theoretical penetration rate calculated from the cutter life in metres and in hours between the actual cutter changes.

I c ,i =

H m ,i H h ,i

(m/h)

[2.6]

Hm,i = calculated cutter life between cutter changes no. i and i+1 (m/cutter) Hh,i = calculated cutter life between cutter changes no. i and i+1 (h/cutter) The ratio Ic /In is used to evaluate the quality of the calculated cutter life. When Ic /In is close to 1.0, the cutter wear situation has been normal. When Ic /In is substantially less or larger than 1.0, it is an indication of: • Abnormal wear of the cutters, e.g. blocked cutter(s). The low cutter life and Ic /In ratio between chainages 179.5 and 181.7 is caused by two blocked cutters being replaced at chainage 181.7, see Table 2.2. • Irregular replacement of cutters, e.g. replacement of one or more positions before the cutter ring is worn to its limit. The low Ic /In ratio between chainages 228.0 and 232.6 is caused by one cutter position being replaced at chainage 232.6 as part of a larger series of neighbouring positions being replaced, see Table 2.2. The ratio Ic /In should be used when deciding suitable rock sampling sites to compare the calculated cutter life to geological parameters. 32



1.6 h/c 1.4 1.2 1.0 0.8 0.6 0.4 0.2 0.0 100

150

200

250

300

350 Chainage, m

Figure 2.4 Instantaneous cutter consumption along the tunnel, based on the data in Table 2.2. Figure 2.4 shows that the representative cutter life for the tunnel section until chainage 300 is at approximately 1.2 – 1.3 h/cutter.

33


2.3 Ring Wear

2.3 RING WEAR Data on the ring wear process is used as a basis to evaluate and improve: • • • •

The ring design and the ring steel The cutterhead design The cutter replacement pattern The boring process.

Wear data are collected at the cutterhead during cutter inspections or cutter changes. The extra few minutes used to collect the wear data from selected cutter positions will give valuable information. Wear Height The wear rate of the individual cutter ring is measured as loss of ring height, see Figure 2.5. The ring height should be measured at the cutterhead a few times during the ring life for selected cutter positions, and for each replaced ring at the cutter repair shop.

h o

h o

h i

h i

h r

h r

0 0

Figure 2.5 Measurements of ring height. h0 = height of new ring (mm) hi = intermediate ring height measured at the cutterhead (mm) hr = ring height at replacement (mm). 34


hwi = h0 − hi hwr = h0 − hr

hwi hwr

2.3 Ring Wear

(mm)

[2.7]

= intermediate wear height at the cutterhead = wear height at time of replacement.

Figure 2.6 shows a typical wear progress of a face cutter ring of the constant cross section type in very hard rock. The wear was measured at each stop for cutter inspection. The wear progress is typical for hard and strong rock, with a high wear rate for new cutter rings (from machine hours 2242.8 to machine hours 2244.6). At machine hours 2244.6 and 2264.1, one neighbour cutter position was replaced, resulting in a lower wear rate until the next cutter change. Of cutter ring properties, the wear rate is mainly dependent on the ring steel quality (e.g. measured as Rockwell C hardness) and the cutter edge width. 35 mm

2283.9

30 2277.1 25 2264.1 20

15 2252.2 10 2244.6 5 2242.8 0 2240

2250

2260

2270

2280

2290

Machine hours, h

Figure 2.6 Wear progress of a cutter ring in a face position. dc = 483 mm. In general, the cutter wear rate increases towards the gauge. Thus, the gauge positions are equipped with wider rings than the face positions. The ring edge width is important to record when the ring wear rate is analysed. 35


2.3 Ring Wear

Figure 2.7 shows average wear height at cutter replacement, measured at the cutter repair shop. The decreasing wear height towards the gauge is important to observe. The main reason for less allowable wear in these positions is the need to maintain the tunnel diameter for geometrical reasons. Furthermore, too large wear height in the gauge will create poor conditions for the new gauge cutter rings when the boring is resumed after the cutter change. Position 20 in Figure 2.7 has very large wear height due to swapping of used cutters from the outer gauge to this position. Hence, the cutters in this position have deliberately been replaced very late. 45 mm 40 35 30 25 20 15 10 5 0 1

2

3

4

5

6

7

8

9

10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 Cutter Position

Figure 2.7 Average cutter ring wear height for the same TBM and tunnel as in Table 2.2. Only cutters replaced due to wear are included. Ring Steel Quality The ring steel quality is measured as Rockwell C hardness (HRC). It is convenient to use a hardness tester like Equotip1 or similar. The chemical composition of the steel is usually not examined at the tunnel site.

1

Manufactured by PROCEQ SA, Switzerland.

36


2.3 Ring Wear

The ring steel hardness should be measured over a cross section of the ring as shown in Figure 2.8. Results of such measurements are shown in Figure 2.9. To do so, the new ring has to be cut, which means that the hardness measurements will be quite expensive and should be limited to a few rings. An alternative method is to measure the hardness at the surface of the ring as shown in Figure 2.8, supplied with measurements over the cross section of the ring for selected ring types. Selected rings are measured as new rings at the cutter shop before they are mounted on the hub and as used rings when the worn ring has been cut and removed from the hub. When the hardness is measured on selected rings on the cutterhead, it is important that the measurements at consecutive cutter inspections or changes are taken at the same positions along the ring. This may be ensured by defining a known point at the hub as starting point and take the measurements at given intervals along the ring. 1 2

3 4

5 M e a s u r e m e n ts o v e r th e c r o s s s e c tio n

6 M e a s u re m e n ts o n th e r in g s u r fa c e

Figure 2.8 Measurement pattern for hardness measurements, with measurement lines and points.

37


2.3 Ring Wear

5 8

a

5 8

c

5 6

c

5 6

b

a d

5 7 e f

b

5 5 5 7 -3 7 3 5

5 7 d

5 5 e

5 4 f

5 3

5 0 4 9 4 6

4 8

4 4

5 0 5 2

5 1

4 9

4 8 4 7

4 6 4 5

4 5

4 4

Figure 2.9 Ring steel hardness (HRC) for two types of rings, measured over the cross section according to Figure 2.8. The distance between measurement points along the measurement lines was 7 mm.

Ring Wear Type

M u s h ro o m

w id th

C h ip p e d a r e a

E d g e w id th

M u s h r o o m in g

C h ip p in g

Figure 2.10 Destructive wear types of cutter rings.

38


2.3 Ring Wear

The ring wear should be abrasive, and not destructive. Destructive wear occurs in two main types for steel ring cutters: • Chipping along the cutter edge • Mushrooming of the cutter edge. Chipping along the cutter edge indicates that the steel hardness (or brittleness) is too high with regard to the cutter thrust and/or the rock strength. Mushrooming of the cutter edge indicates that the steel hardness (or brittleness) is too low with regard to the cutter thrust and/or the rock strength.

Figure 2.11 Chipping of a cutter ring.

The chipping or mushrooming of the cutter edge is quantified as: • Percentage of the ring circumference affected by the destructive wear • The extra edge width due to mushrooming • The typical length of individual chippings. The measurements should apply to each side of the ring edge, denoted as inner and outer side in relation to the cutterhead centre. 39

3. PENETRATION TESTS 3.1

3.1 Test Procedure

TEST PROCEDURE The main purpose of a penetration test is to evaluate the machine performance in a given geology. Hence, a penetration test should be followed by a detailed engineering geological mapping of the actual tunnel section. In addition, the rock drillability should be tested in the laboratory. A complete penetration test consists of the following parts: • Measurement of the cutterhead penetration over a given time at various thrust levels and constant RPM. • Registration of the average cutterhead torque of each cutter load level. • Registration of other relevant data such as cutter wear state, whether the test is at the start, middle or end of the stroke, cutterhead vibration level, etc. • Measurement or registration of the net penetration rate, cutter thrust level and cutterhead torque of the previous and following strokes. • Collecting a complete chip sample for the penetration test and chip samples for the previous and following strokes.

Thrust levels A penetration test should include at least four thrust levels. When deciding the thrust levels Mt to be used, the current thrust level used by the operator is selected as the 100 % level, denoted as MB100. This presupposes that the machine is operated at an optimum thrust level concerning net penetration rate, cutter life, cutterhead vibrations, etc. The thrust levels of the test are then selected as: M t1 ≈ 0.7 ⋅ M B100 M t 2 ≈ 0.8 ⋅ M B100 M t 3 ≈ 0.9 ⋅ M B100

(kN/cutter)

[3.1]

M t 4 ≈ 1.0 ⋅ M B100

If applicable with regard to the cutter life, available torque, etc., it is recommended to include a fifth thrust level in the test:

40

3. PENETRATION TESTS

M t 5 ≈ 1.05 ⋅ M B100

(kN/cutter)

3.1 Test Procedure

[3.2]

The actual thrust levels are selected as rounded numbers to ease the operation of the TBM during the test. If the machine in the given rock conditions is operated at a thrust level corresponding to a cylinder pressure of pB100 = 217 bar, the actual thrust levels are selected as follows: p t1 = 160 p t 2 = 180 p t 3 = 200

(bar)

[3.3]

p t 4 = 220 p t 5 = 230

To ensure the quality of the test, the operator must maintain a constant thrust level during each step of the test. The applied thrust level must be recorded for each step of the test, either as an observed average from the thrust cylinder pressure gauge or as a printout from the onboard computer. The applied thrust must be averaged over time.

Test Duration The penetration is measured over a time tt corresponding to approximately 30 revolutions of the cutterhead at each thrust level. For small machine diameters, the penetration should be measured over at least tt = 3 minutes. Before the measurement of the penetration starts, the operator must stabilise the thrust at each level.

Penetration Measurement The penetration it is measured in mm over the given time. It is recommended that the penetration measurements are taken at one of the thrust cylinders since these are in

41


3.1 Test Procedure

direct contact with the cutterhead. One may also use instruments at the operator console when the accuracy of such instruments has been verified.

Torque Measurement The applied torque is recorded as the average amperage It for each step of the test, either as an observed average of the ampere meters or as a printout from the onboard computer. The applied amperage must be averaged over time. The applied voltage of the cutterhead drive motors must be checked. The motors should be operated at the rated voltage UN. Otherwise, the applied voltage of the test Ut must be noted.

Test Data Example Tunnel:

Chainage:

Date:

Average Applied Pressure pt (psi)

Average Applied Amperage It (A)

Penetration it for tt = 3 min (mm)

2400

135

70

2700

175

97

3000

225

132

3300

250

169

RPM: 13.4

Table 3.1

Comments: Average cutter wear. The test started right after regripping.

Example of data recorded during a penetration test. For further treatment of the test, see Section 3.2

Previous and Following Strokes It is recommended to measure the net penetration rate of the previous and following strokes of the test stroke. This is done to ensure the quality and representativeness of the test. The penetration rate should be measured over the complete strokes. For large 42


3.1 Test Procedure

TBM diameters one may measure the net penetration rate only over parts of the stroke length, e.g. for 20 - 30 minutes.

Personnel At least two persons in addition to the TBM operator are needed to perform a penetration test with chip sampling: One for the penetration measurements and one for the chip sampling.

43



3.2 PENETRATION CURVE The penetration curve in [3.4] is found by treatment of the observed data as shown below.

b

M  i 0 =  t  (mm/rev)  M1 

i0 Mt M1 b

[3.4]

= basic penetration (mm/rev) = gross average thrust of each thrust level (kN/cutter) = critical thrust to achieve a penetration of 1.0 mm/rev (kN/cutter) = penetration coefficient.

The data from Table 3.1 are modified to the data given in Table 3.2. The actual machine has Ntbm = 25 cutters and ntc = 2 thrust cylinders, each with an effective diameter of dtc = 444.5 mm. The applied gross average thrust is calculated from the applied cylinder pressure according to [3.5]. The basic penetration is found by [3.6].

π ⋅ d tc2 0.4536 ⋅ g M t = p t ⋅ ntc ⋅ ⋅ 2 4 ⋅ 25.4 1000 ⋅ N tbm

( p t in psi)

p π ⋅ d tc2 1 M t = t ⋅ ntc ⋅ ⋅ 10 4 1000 ⋅ N tbm

( p t in bar)

pt ntc dtc g Ntbm

i0 =

(kN/cutter)

[3.5]

= applied thrust cylinder pressure (psi or bar) = number of thrust cylinders = effective diameter of the thrust cylinders (mm) = gravitational constant = 9.81 (m/s2) = number of cutters on the cutterhead.

it RPM ⋅ t t

(mm/rev)

[3.6]

44



i0 = basic penetration (mm/rev) it = penetration over the time tt (mm) RPM = cutterhead rotation speed (rev/min) tt = test duration time for one thrust level (min) Gross Thrust Mt (kN/cutter)

Basic Penetration i0 (mm/rev)

log10 Mt

log10 i0

205.5

1.74

2.3128

0.2405

231.2

2.41

2.3640

0.3820

256.9

3.28

2.4098

0.5159

282.6

4.20

2.4512

0.6232

Table 3.2

Cutter thrust and basic penetration for a penetration test.

M1 and b in [3.4] are found by linear regression of the log10 values of the thrust and the penetration. It is convenient to use a spreadsheet to perform the necessary calculations. An example of an EXCEL spreadsheet for this purpose is shown in appendix E. The log10 values of Mt and i0 will usually fit very good to a straight line, see data of Test 1 in Figure 3.1. The xy-plot of the log10 values should be analysed with regard to which observations to include in the regression and the penetration curve. Test 2 is an example of disturbed data, in this case a Marked Single Joint was found in the consecutive mapping of the actual tunnel section. Another cause of disturbance is too large variation of the thrust during one or more steps of a test. When data of one or two thrust levels are disturbed, the test may still be used for analyses, but with caution. The linear regression of the log10 data results in the equation in [3.7]. The log10 version of [3.4] is shown in [3.8]. The constants of the linear equation are transformed to M1 and b according to [3.9] and [3.10]. log 10 (i 0 ) = AR ⋅ log 10 ( M t ) + B R

AR BR

[3.7]

= regression constant, see Appendix E = regression constant, see Appendix E. 45



log 10 (i 0 ) = b ⋅ (log 10 ( M t ) − log 10 ( M 1 ))

[3.8]

M1 is defined as the critical thrust to result in a penetration of 1.0 mm/rev. Hence, M1 is found by solving [3.7] for i0 = 1 mm/rev (log10(1) = 0).

M 1 = 10

−

BR AR

(kN/cutter)

[3.9]

b is found by setting [3.7] = [3.8] and substituting M1 with [3.9].

b ⋅ (log 10 ( M B ) − log 10 (10

−

BR AR

)) = AR ⋅ log 10 ( M B ) + B R

[3.10]

⇒ b = AR

1 log10 i0

Test 1 Test 2 Linear regression of Test 1 Linear regression of Test 2

0.8

0.6

0.4

0.2

0 2.3

2.35

2.4

2.45

2.5 log10 M t

Figure 3.1 Plot of log10 values of Mt and i0 for two penetration tests, including linear regression curves.

46



3.3 CUTTER COEFFICIENT The calculation of the cutter coefficient kc and the cutter constant cc is based on the registrations of the applied amperage during the penetration test, see Table 3.1. The use of electric motors is presupposed. To be able to calculate kc, one needs the rated amperage IN and the efficiency curve (cosφ⋅η) of the motors. An example of this is shown in Table 3.3 and Figure 3.2. Applied Amperage It (A)

cosφ

η

It /IN

Efficiency cosφ⋅η

54

0.524

0.931

0.153

0.479

100

0.717

0.944

0.284

0.677

160

0.850

0.960

0.455

0.816

200

0.869

0.963

0.568

0.837

300

0.876

0.961

0.853

0.842

400

0.852

0.957

1.137

0.815

Table 3.3

Efficiency of an electric cutterhead drive motor with a rated power of PN = 336 kW and a rated amperage of IN = 351.8 A. The necessary data are usually found in the Service and Operation Manual of the machine.

The cutter coefficient is calculated according to [3.11].

kc =

U t ⋅ I t ⋅ 3 ⋅ (cos φ ⋅ η ) ⋅ 60 ⋅ n m ⋅ 2 2 ⋅ π ⋅ d tbm ⋅ rmc ⋅ M t ⋅ 1000 ⋅ N tbm ⋅ RPM

[3.11]

Ut = applied voltage of the cutterhead drive motors (V) It = applied amperage of the thrust level (A) nm = number of motors on the cutterhead dtbm = TBM diameter (m) rmc = relative radius to the position of the average cutter position, see [3.12] Mt = applied thrust (kN/cutter) Ntbm = number of cutters on the cutterhead RPM = cutterhead revolutions (rev/min) 47



1.0 cosφη 0.9

0.8

0.7

0.6

0.5

0.4 0.0

0.2

0.4

0.6

0.8

1.0

1.2 It/IN

Figure 3.2 Efficiency curve of the electric motors in Table 3.3.

N tbm

∑r i =1

i

rmc =

N tbm 0.5 ⋅ d tbm

ri

= radius to position of cutter no. i (m)

[3.12]

The cutter constant is calculated according to [3.13]. The resulting cutter constant of a penetration test is shown in Figure 3.3.

cc =

kc

i0

= basic penetration of the thrust level (mm/rev)

i0

[3.13]

48



0.050 cc 0.045

0.040

0.035

0.030 200

220

240

260

280

300 Mt, kN/cutter

Figure 3.3 Cutter constant of the penetration test in the Tables 3.1 and 3.2. Ut = 660 V, nm = 4, dtbm = 3.5 m, dc = 483 mm, rmc = 0.6334, Ntbm = 25.

Figure 3.3 shows an irregular progress of the cutter constant with regard to increased thrust level. One reason may be incorrect reading of the Ampere meters, but most likely it is because the two lowest thrust levels have a quite low basic penetration, see Table 3.2, making [3.13] a somewhat incorrect approximation.

49


3.4 Specific Energy

3.4 SPECIFIC ENERGY The gross specific energy used to break the rock may be estimated from the penetration test data or from shift log data. The specific energy estimated on that basis does not include the cutterhead thrust system which usually has an installed power of 10 % or less compared to the cutterhead torque system. Since the cutterhead torque is used to estimate the specific energy for TBM boring, the muck removal through the cutterhead is included. The contribution of the muck removal is negligible compared to the contribution from the cutterhead rotation. The specific energy consumption is calculated according to [3.14] and plotted as in Figure 3.4.

Ws =

U t ⋅ I t ⋅ 3 ⋅ (cos φ ⋅ η ) ⋅ n m ⋅ 3600 ⋅ 1000 ⋅ 4 2 i 0 ⋅ RPM ⋅ 60 ⋅ π ⋅ d tbm ⋅ 1000 2

(MJ/m3)

[3.14]

150 MJ/m3 140

130

120

110

100 200

220

240

260

280

300 Mt, kN/cutter

Figure 3.4 Specific energy consumption based on data from the penetration test in Tables 3.1 and 3.2.

50

4. CHIP ANALYSES 4.1

4.1 Test Procedure

TEST PROCEDURE Analyses of the largest chips produced by a TBM may give information of the boring process and the rock breaking mechanisms, as well as material properties of the TBM muck and drillability parameters of the intact rock. Chip sampling must be combined with registration of concurrent machine and performance data. Chip sampling is of extra value when combined with a penetration test. The following sampling procedure is related to penetration tests, but may be used as a general procedure when sampling chips. A chip sampling test should be followed by a detailed engineering geological mapping of the actual tunnel section. During a penetration test, the largest chips are sampled, measured and analysed as described below.

Sample Site A chip sample should be collected as close to the cutterhead as possible. At an open hard rock TBM this means at the machine conveyor or at the bridge conveyor close to the rear end of the TBM. Caution must be exercised when the chips are picked from the conveyor belt.

Sample Time When the operator has stabilised the thrust level of the machine, the penetration measurement starts. After that, one should wait at least half a minute before the chip sampling starts, to ensure that the chips sampled are inside the penetration measurement. One should stop the sampling when the penetration measurement stops.

Sample Size A chip sample should consist of 20 large chips at each thrust level. This is achieved by picking a larger number of chips (25 - 30), sorting the chips according to size by 51

4. CHIP ANALYSES

4.1 Test Procedure

visual judgement, and then discarding the smallest, leaving the 20 largest chips for measurements and possible laboratory testing. The number of chips is mainly decided by the time available for sampling at each thrust level and by the necessary number of chips to establish a stable mean value and standard deviation of the chip sizes. Figure 4.1 shows the fluctuation in the mean value and the standard deviation of the thickness of the chip sample in Table 4.1.

30 mm 25 Mean value 20

15

10

5 Standard deviation 0 0

5

10

15

20 25 Number of chips measured

Figure 4.1 Mean value and standard deviation of the chip thickness as a function of number of chips in the sample.

Normally, the largest chips will be produced by the face cutters. Chips from the gauge will usually not be picked since they are thinner and less wide than chips from the face cutters. Chips produced by the centre cutters will be recognised by the evident ring shape. Chips that by visual judgement originate from the centre or gauge should be discarded from the sample. Chips that seem to be broken during removal from the rock face to the sampling site are discarded. Blocks that are fall-outs from Marked Single Joints or similar are also discarded. 52

4. CHIP ANALYSES

4.1 Test Procedure

Muck Sample When collecting a muck sample for sieve testing, the sample volume should be at least 20 l and may be collected at conveyor transfer points to get the best representative sample. Muck samples collected from e.g. a truck load or a muck pile may be subject to separation of the fines from the coarser chips, and therefore not representative samples.

53

4. CHIP ANALYSES

4.2 Chip Size

4.2 CHIP SIZE The size of each chip is measured as largest length, width and thickness, regardless of where along the chip the largest size occurs. It is of course important that the three measurements are taken more or less perpendicularly to each other. The chip size measurements are averaged for each thrust level as in the Tables 4.1 and 4.2, and plotted as a function of thrust as in Figure 4.2. The cubic chip size Vch is calculated by [4.1] and plotted as in Figure 4.3. The chipping frequency fch is calculated by [4.2] and plotted as in Figure 4.4. Chip no.

Height hch (mm)

Width wch (mm)

Length lch (mm)

1

24

65

267

2

43

65

209

19

16

67

193

20

26

66

204

Mean size (mm)

28.5

66.6

217.2

Standard Deviation, mm

7.9

5.8

37.2

Standard Deviation, %

27.7

8.7

15.4

⋅⋅⋅⋅⋅

Table 4.1

Chip size measurements for one thrust level of a penetration test. dtbm = 8.5 m, dc = 432 mm. The rock type is mica gneiss. Average Chip Size (mm)

Thrust Mt (kN/cutter)

Penetration i0 (mm/rev)

hch

wch

lch

Chipping Frequency fch (rev-1)

173.4

2.00

26.8

55.8

215.5

0.075

183.1

2.20

24.9

64.4

212.0

0.088

192.7

2.51

22.3

62.9

205.6

0.112

202.3

2.58

25.4

66.1

233.7

0.101

221.6

3.41

28.5

66.6

217.2

0.119

Table 4.2

Chip size measurements and chipping frequency of a penetration test. Data of the same test as in Table 4.1. 54

4. CHIP ANALYSES

4.2 Chip Size

300 mm

Thickness Width Length

250

200

150

100

50

0 150

175

200

225

250 Mt, kN/cutter

Figure 4.2 Average size of the largest chips from the penetration test in Table 4.1.

500 1000 mm3 450

400

350

300

250 150

175

200

225

250 Mt, kN/cutter

Figure 4.3 Cubic chip size of the penetration test in Table 4.2. 55

4. CHIP ANALYSES

4.2 Chip Size

Vch = l ch ⋅ wch ⋅ hch

lch wch hch

f ch =

i0

(mm3)

[4.1]

= average length of largest chips (mm) = average width of largest chips (mm) = average thickness of largest chips (mm)

1 hch i0

(rev −1 )

[4.2]

= TBM penetration per revolution (mm/rev)

0.14 -1

rev

0.12

0.1

0.08

0.06 150

175

200

225

250 Mt, kN/cutter

Figure 4.4 Chipping frequency of the penetration test in Table 4.2.

56

4. CHIP ANALYSES

4.2 Chip Size

The shape factor is a relative measure of the shape of an individual chip or of the average dimensions of 20 large chips. The calculation of the shape factor is given in [4.3]. A plot of the shape factor of the chips from the penetration test in Table 4.2 is shown in Figure 4.5. f hw = hch / wch

[4.3]

f wl = wch / l ch

1 fwl 0.9

FLAT

CUBIC

0.8 0.7 0.6 0.5

ELONGATED 0.4 0.3 0.2 0.1 0 0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1 ftw

Figure 4.5 Chip shape of the average chip size of the penetration test in Table 4.2.

57

4. CHIP ANALYSES

4.3 Sieve Curve

4.3 SIEVE CURVE The sieve curve is found by sieving the TBM muck at square sieves. The sieve sizes to find a rough particle size distribution of the TBM muck are recommended as: • 64, 45, 32, 16, 8, 4, 2, and 1 mm. The sieve curve established by the above sieve range should only be used for evaluation of the rock breaking and boring process. To perform a standard sieve test to characterise the material for technical or building purposes, the finer fractions must be decided by sieves down to 0.063 mm and by sedimentation for the even finer fractions. Using a sieve shaker will give a good result for the fractions below 2 mm. Experience shows that from the 4 mm sieve and upwards, some chips that should have passed will be left at the sieve due to the flat and elongated chip shape. Hence, the sieve test must include a manual verification of the size of each chip as illustrated in Figure 4.6.

d s d s

Figure 4.6 Manual sieving of TBM chips. ds = square sieve size.

A chart for presentation of the sieve test results is shown in Appendix D. Normalised sieve curves are found in the Project Report 1F-98 HARD ROCK TUNNEL BORING The Boring Process.

58

4. CHIP ANALYSES


4.4 KERF DEPTH FACTOR The kerf depth factor is measured where the chip has its largest thickness. In the laboratory, the chips are cut by diamond saw, which gives the best measurements. In the field one has to rely more on judgement as the measurements are taken on uncut chips. Concerning the accuracy of the measurements, one should bear in mind that the kerf depth factor is used to show a trend, not for precise calculations. Furthermore, when the number of chips increases, the influence of the random measurement error is decreased. The kerf depth is measured as shown in Figure 4.7. The kerf depth factor is:

f kd =

ik hch

fkd ik hch

= kerf depth factor = kerf depth at chip forming (mm) = maximum chip thickness (mm)

ik

[4.4]

C u tte r e d g e in d e n ta tio n h

c h

Figure 4.7 Measurement of the kerf depth at chip forming.

59

4. CHIP ANALYSES


It may be difficult or uncertain to identify the exact kerf depth or the cutter ring indentation of each chip. The best indicator is the fine powder that may be found at the side of the chip, combined with the form of the chip side towards the kerf, see Figure 4.8. Deficient identification of the kerf depth should lead to discarding the chip from the kerf depth measurements.

60

4. CHIP ANALYSES

4.5 Crack Growth

4.5 CRACK GROWTH The best approach to study the crack growth and the chipping mechanisms for hard rock tunnel boring is to examine cross sections of the large chips. When the chips are cut by diamond saw, the cracks inside the chip are easy to identify. The chip forming cracks are of course found as the surface of the chip. Usually, the chip is cut where its thickness is largest. It is supposed that the final forming of the chip occurred at that part of the chip. A three-dimensional model of the crack pattern may be established by cutting the chip in equally spaced slices. The orientation of the chip with regard to which side has been towards the cutterhead is decided by the chip form and by the fine powder remaining in the kerfs made by the cutter along the chip. Figure 4.8 shows a cross section of a chip with easily identifiable cutter tracks (kerfs) and radial cracks from cutter passes prior to the chip forming.

C ra c k s fro m

h

p r e v io u s p a s s e s

ik c h

ik h

c h

Figure 4.8 Identification of cutter tracks and radial cracks in a chip.

Crack growth is best observed in isotropic and fine-grained rock types.

61

4. CHIP ANALYSES

4.6 Rock Properties

4.6 ROCK PROPERTIES The largest chips may be used for testing of drillability, strength and wear properties. Actual tests to be performed on chips are DRI, Mini-DRI, CLI, CAI and Point Load Strength. One must however observe that the chips have been subject to high stress that may have influenced the rock properties somewhat.

DRI The Brittleness Value S20 of the DRI test is believed to be influenced by the type of test material. TBM chips will most likely contain some small cracks from the boring process. Parallel tests of 50 mm diameter cores and TBM chips indicate that the Brittleness Value (and therefore the DRI value) tested on chips is 2 - 4 units higher than the Brittleness Value tested on 50 mm cores. The few available data covers the Brittleness Value range of 35 - 50. Since the systematic error in the DRI value tested on chips seems to be relatively small, it is advisable to use TBM chips as test material in the DRI test. The chips must be of a size and orientation so that one is able to perform the Sievers' J-value (SJ) test parallel to the possible foliation of the rock. This is achieved in most cases by cutting the chips in the desired orientation. The SJ test may be influenced by microcracks resulting from the cutter forces.

Mini-DRI The brittleness test of the Mini-DRI uses the crushed aggregate fraction of 2 - 4 mm. The Mini-DRI value is believed to be little influenced by the type of test material as long as it is chips, cores or blocks. Hence, the largest TBM chips are well suited to be used in the Mini-DRI test. The SJ test requires the same size and orientation of the chips as the standard DRI test.

CLI The abrasion tests AV and AVS use the crushed fraction of grain size < 1 mm. The AV and AVS value are not believed to be influenced by the type of test material as long as it is crushed from larger pieces, and the largest TBM chips are well suited for these tests. In order to calculate the CLI, the chips must be of a size and orientation so that one is able to perform the SJ test parallel to the possible foliation of the rock. 62

4. CHIP ANALYSES

4.6 Rock Properties

CAI The CAI test is a scratch test performed on a rough rock surface. The TBM chips are well suited as test material to test the CAI value.

Point Load Strength The largest chips may be suited for point load testing, providing that the chips selected are oriented parallel or perpendicular to the rock structure. The point load test may be carried out in the field by portable equipment, giving an immediate measure of the rock strength. A correlation between the point load index Is and the DRI is found in the Project Report 1E-98 HARD ROCK TUNNEL BORING Geology and Site Investigations. The Point Load test may be influenced by microcracks resulting from the rock cutting process.

63

5. BACK-MAPPING

5.0 Introduction

5.0 INTRODUCTION The engineering geological back-mapping of a bored tunnel should consist of the following: • Continuous and detailed mapping of rock mass fracturing • Continuous and detailed mapping of rock type distribution • Rock sampling and laboratory testing of rock properties. The engineering geological back-mapping establishes a geological model of the tunnel, to be used for evaluation of the machine performance, cutter life, machine utilisation, etc. It is convenient to map longer sections, e.g. 500 m, for each mapping round. Backmapping in tunnels lined with concrete elements or shotcrete at the cutterhead is very difficult to perform with the purpose to establish a continuous geological model. In such tunnels the back-mapping must be improvised and done at points whenever the rock surface is available. The mapping is a subjective task with regard to degree of fracturing, type of fractures, rock type, etc. The quality of the mapping will improve substantially when a team of two persons is doing the back-mapping. The parameters used in the back-mapping are the same as those used in the site investigations, as described in the Project Report 1E-98 HARD ROCK TUNNEL BORING Geology and Site Investigations.

64

5. BACK-MAPPING


5.1 MAPPING PROCEDURES The mapping in the tunnel should be recorded on a sheet with standardised entries. An example of a completed sheet for a 50 m tunnel section is shown in Figure 5.1. A blank form is found in Appendix D, and is also available in digital format1. T U N N E L : M E R K R A F T C h a in a g e 4 5 5 0

S ig n a tu r e : A B / B S

D a te : 3 / 2 - 9 2

7 0

8 0

9 0

4 6 0 0

L e ft w a ll

6 0

1 3 5 ° 9 0 ° 4 5 °

R o o f

1 0 ° 1

R ig h t w a ll

-4 5 ° -9 0 ° -1 3 5 °

G r e e n s c h is t (G r e e n s t o n e )

C o m m e n ts

F r a c tu r in g

R o c k ty p e

12 g / 3 9 °

1

S t III ÷

D r ip p in g

S t II - III

f o l: 2 1 1 g / 4 2 ° tu n n e l ~ 2 8 0

S t II +

a @ 3 6 °

S t III 1

S t III

C la y f ille d 1 - 2 c m

g

Figure 5.1 Completed back-mapping sheet from a 3.5 m diameter tunnel.

1

Contact the author.

65

5. BACK-MAPPING

chainage


+

+

+

+

+

+

left wall

roof

right wall

left wall

roof

right wall

0°

0° 45°

roof

roof

-45° 60°

90°

left wall

right wall

135°

-90°

-60°

90°

-90° left wall

right wall

-135° 180° -180°

180° -180°

dtbm < 5 - 8m

dtbm > 5 - 8m

Figure 5.2 Representation of the tunnel circumference in the mapping sheet. Right and left wall applies when looking towards the tunnel face.

The mapping is concentrated on 10 m sections. Each section is mapped according to the following steps: • Record the Marked Single Joints and measure strike and dip where convenient. • Record other singular phenomena like intrusions, water, rock fall-outs, rock support, etc. • Decide the rock type. • Decide the number of fracturing systems (usually one or two). • Decide the type of fracturing for each system (St or Sp). • Decide the degree of fracturing for each system. • Measure strike (αs) and dip (αf) of the fracturing system(s) at least once for each 50 m. • Measure the strike of the tunnel (αt) at the same time as the strike and dip of the fracturing is measured. The registrations should be an average for the 10 m section. If necessary (e.g. change of rock type), the 10 m section may be subdivided.

66

5. BACK-MAPPING


The degree of fracturing should consequently be evaluated along a scanline at one of the tunnel walls, preferably in the lower part of the wall since that gives a good possibility to take a closer look at the rock mass. In special cases when the scanline along the wall is obviously not representative for the tunnel section (e.g. for intrusions) one may evaluate the section volume as a whole. In most cases, the degree of fracturing is recorded by visual interpretation with occasional measurements of the distance between the planes of weakness. The distance is measured perpendicular to the planes of weakness as shown in Figure 5.3. = fis s u r e s -4 5 ° R ig h t tu n n e l w a ll

2 4 4 0

S c a n lin e fo r m a p p in g -1 3 5 °

0 1 3 5

a f

=

2 4 3 5

6 7 5

0 1 2 5

9 0 0

(1 3 5 0 + 6 7 5 + 1 2 5 0 + 9 0 0 + 1 0 0 0 )m m 5

0 0 1 0

= 1 0 3 5 m m

» S t I-

Figure 5.3 Measurement of fracture spacing. The fissures have strike approximately perpendicular to the tunnel axis.

Figure 5.4 shows two examples where it may be difficult to evaluate the degree of fracturing. Case 1 has various possibilities, of which 3 are evaluated below. • The rock mass is massive (homogeneous) outside the tunnel. Hence, there are only 3 fissures to consider, and the average distance between the fractures may be estimated as:

af =

(500 + 600) mm = 550 mm 2

67

5. BACK-MAPPING


This is obviously ambiguous, and other possibilities must be considered. • The "next" fissures are just outside the tunnel in the roof and the invert. Average distance between the fractures will be:

af =

(800 + 500 + 600 + 1600) mm = 875 mm 4

875 mm corresponds approximately Fissure Class I- (St I-). • The "next" fissures are 800 mm outside the tunnel roof and 1600 mm outside the tunnel invert. Average distance between the fractures will be:

af =

(800 + 800 + 500 + 600 + 1600 + 1600) mm = 1475 mm 4

1475 mm corresponds approximately Fissure Class 0-I (St 0-I). The conclusion for Case 1 is that the degree of fracturing may be classified as St Ior lower, with St 0-I as the recommended classification. Case 2 also has various possibilities, of which 2 are evaluated below. • The rock mass is massive (homogeneous) outside the tunnel. Hence, there are 6 fissures to consider, and the average distance between the fractures may be estimated as:

af =

500 mm = 100 mm 5

This is obviously ambiguous, and other possibilities must be considered.

68

5. BACK-MAPPING


• If the "concentration" of fissures or joints has some extent along the tunnel (e.g. more than 2 ⋅ dtbm), it may be regarded as a weakness zone and classified as a prominent Marked Single Joint. The conclusion for Case 2 is that the concentrated fissures or joints may be classified as a Marked Single Joint. It is however difficult to give general rules for similar cases for when to classify as Marked Single Joints and when to classify as systematic fractured rock mass.

e s s u r F is

m m 8 0 0

m 0 m 2 2 0

m m 5 0 0

m m 6 0 0

m m 5 0 0

m 0 m 1 6 0

C a s e 1

m m 8 0 0

re s s s u 6 fi

C a s e 2

Figure 5.4 Examples of evaluation of the degree of fracturing in a 3.5 m diameter tunnel. The fissures have strike approximately perpendicular to the tunnel axis and the dip is close to horizontal.

The best way to adjust and fine tune the classification of the degree of fracturing as seen in the tunnel wall, is to observe the tunnel face during a standstill of the TBM and experience how the fracturing is utilised in the rock breaking process. The notation and registrations of the back-mapping may be individual, but some basic rules should be applied: • The preferred sectioning of the tunnel circumference is as follows: The roof and the walls of the mapping form cover a sector of 90° each, leaving a sector of 90° in the invert. When necessary, the chainage field and the field below the right wall field are utilised to record Marked Single Joints etc., see Figure 5.1. 69

5. BACK-MAPPING


• For large diameter tunnels, the roof and wall sectors of the mapping form may be increased to 120° to be able to cover the complete tunnel (see Figure 5.2 and Appendix D). • Marked Single Joints are recorded as continuous lines, other rock mass fracturing is recorded as dotted lines. • The type and degree of fracturing must be noted for each 10 m section. Do not use arrows or similar to indicate "continues from, or the same as, the previous section". • Strike and dip is measured in a 400g or 360° scale, with the dip direction always to the same side of the strike direction2.

Quality Control The engineering geological back-mapping is a highly subjective process, with a strong element of judgment in the classification of the degree of fracturing. To ensure a good quality of the classification, the following procedures are recommended: • A team of two persons should perform the back-mapping. • Evaluate each 10 m section individually, and use enough time when mapping. • Before the mapping of a new tunnel section starts, check and evaluate the mapping of the last 30 - 50 m of the previous section. This will adapt the mapping personnel to the site geology before mapping of the actual section starts and uncover possible disagreement in the classification of the previous section. • When the mapping of a tunnel section is finished, re-map selected subsections of 20 - 30 m length at 150 - 200 m intervals. Using only 10 m subsections when remapping will not give the desired quality control. • When the mapping is finished, check the mapping of selected subsections against the machine performance to confirm that the variation in the machine performance corresponds to the variation in the degree of fracturing.

2

See Section 2.1 in the Project Report 1D-98 HARD ROCK TUNNEL BORING Geology and Site

Investigations.

70

5. BACK-MAPPING


5.2 AGGREGATION OF MAPPING DATA The data from the back-mapping must be aggregated to be used further. The geological model of the tunnel is typically divided into sections, in which as many as possible of the boring parameters are constant or showing little variation. The most important parameters to consider when the tunnel is divided into sections, are: • • • • •

Rock type and rock properties such as DRI, CLI, quartz content General level of rock mass degree of fracturing General angle between the tunnel axis and the planes of weakness Tunnel direction Machine parameters such as thrust level, cutterhead RPM, cutter type, etc.

The aggregation of geological data is basically an averaging of rock properties and rock mass degree of fracturing.

Rock Properties If test results for more than one rock sample is available for one geological tunnel section, the arithmetic mean of the laboratory test results or indices are used. In special cases, one may use a weighted average (parameters weighted by tunnel length). When rock samples are assumed to represent a given length of a tunnel section, it is recommended to subdivide the tunnel section rather than to use a weighted average for the original section.

Rock Mass Fracturing The systematic rock mass fracturing of the tunnel section is grouped according to the degree of fracturing as shown in Figure 5.6. The angle between the tunnel axis and the planes of weakness is estimated from pole plots of the strike and dip measurements (see Project Report 1D-98 HARD ROCK

71

5. BACK-MAPPING


TUNNEL BORING Geology and Site Investigations), or calculated as the average of direct measurements or estimates of α taken in the section. The average ks factor of the section is calculated according to [5.1] and Table 5.1.

IV 4 .0 F is s u r e C la s s k

J o in t C la s s

s

3 .0

III-IV

2 .0 III II-III 1 .0

2 0

3 0

4 0

5 0

6 0

7 0

8 0

I-II

II I

I

0 -I

0 1 0

II

0

9 0 a , d e g re e s

Figure 5.5 Fracturing factor. From the Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear. The equations of the ks factor are available from the author. The ks factor of fracturing classes not shown as curves in Figure 5.5 is found by linear interpolation between the given curves.

72

5. BACK-MAPPING


90 m 80 70 60 50 40 30 20 10 0 st I+

st I-II

st II-

st II

st II+

st II-III Degree of fracturing

Figure 5.6 Example of distribution of rock mass fracturing of a 235 m long tunnel section.

n

k s − avg =

li ks-i

∑l

i =1 n

i

li ∑ i =1 k s −i

[5.1]

= tunnel length of fracturing class i of the section = ks factor of fracturing class i of the section.

The ks-avg relates to the basic net penetration rate I0 of the tunnel section, see below.

73

5. BACK-MAPPING


ks-i

li k s −i

∑l

20

0.87

22.99

20

22.99

St I-II

50

0.96

52.08

70

75.07

3

St II-

80

1.06

75.47

150

150.54

4

St II

54

1.15

46.96

204

197.50

5

St II+

20

1.26

15.87

224

213.37

6

St II-III

10

1.37

7.30

234

220.67

i

Fracturing Class

1

St I+

2

Tunnel Length of Fracturing Class i (m)

li

i

∑k

s −i

6

k s −avg =

∑l i =1

i

6

li ∑ i =1 k s − i

=

234 = 1.06 220.67

[5.1]

For α = 25°, ks-avg =1.06 means a Fissure Class of approximately St II-.

Table 5.1

Calculation of ks-avg.

The total tunnel length of the section influenced by Marked Single Joints is found by summing the length in the tunnel of each MSJ. Table 5.2 shows how the 8 Marked Single Joints in Figure 5.1 are treated. According to Figure 5.7, the Marked Single Joints should be grouped in two groups: • MSJs giving a correction factor kesp < 1.4 • MSJs giving a correction factor kesp = 1.4 (the recommended maximum value) This may roughly be achieved by grouping the Marked Single Joints in two categories: • MSJs each influencing a tunnel length less than the tunnel diameter, i.e. αesp > 45° • MSJs each influencing a tunnel length greater than the tunnel diameter, i.e. αesp ≤ 45° The average angle αesp between the tunnel axis and each MSJ group is found by [5.2]. The basic net penetration rate I0 of the tunnel section is found by [5.3]. 74

5. BACK-MAPPING

α esp = arctan

nesp ⋅ d tbm l esp


(°)

[5.2]

nesp = number of MSJs in the MSJ group of the section lesp = cumulated length of MSJs in an MSJ group of the section (m) dtbm = TBM diameter (m)

I0 =

In lj lesp-i kesp-i

l esp −i In (l j − ∑ l esp −i + ∑ ) lj k esp −i

(m/h)

[5.3]

= net penetration rate of the tunnel section (m/h) = tunnel length of the section (m) = cumulated tunnel length of MSJ group i of the section (m) = correction factor MSJ group i of the section, see Figure 5.6.

DRI=30

1.4

DRI=40 DRI=50

kesp 1.3

DRI=60 1.2

1.1

10

20

30

40

50 a

60 esp

, degrees

Figure 5.7 Correction factor for Marked Single Joints. From the Project Report 1B98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear. 75

5. BACK-MAPPING


Depending on the practice used in the back-mapping, it may be necessary to correct the recorded length of Marked Single Joints. When the invert sector of the tunnel is not included in the mapping form (see Figure 5.2), non-vertical joints will not be recorded with their real length in the form. Joints with strike perpendicular to the tunnel axis will have the largest deviation, and the recorded length of these is 75 % of the real influence length. For other strike and dip angles, the correction is estimated roughly. When the mapping form does not include the invert sector, it is a good possibility to use the chainage field and the field below the right wall field to represent half of the invert sector each, as indicated in Figure 5.2. This is utilised for Marked Single Joint no. 2 in Figure 5.1.

MSJ No.

Recorded

Real Length in Tunnel

Cumulated Length in Tunnel

Length in

(m)

(m)

Tunnel (m)

MSJ Group 1

MSJ Group 2

MSJ Group 1 MSJ Group 2

1

8

2

1.5

3

4

5

13

4

4.5

4.5

17.5

5

3.5

3.5

5.5

6

2

2

7.5

7

4.5

6

23.5

8

4

5.5

29

α esp −1 = arctan α esp −2 = arctan

Table 5.2

8 2

n esp −1 ⋅ d tbm l esp −1 n esp −2 ⋅ d tbm l esp − 2

8 2

= arctan

3 ⋅ 3.5 = 54.5° 7 .5

⇒ kesp-1 = 1.4

= arctan

5 ⋅ 3 .5 = 31° 29

⇒ kesp-2 = 1.29

Calculation of lesp and αesp for a 3.5 m diameter tunnel. Data from Figure 5.1. Real Length in Tunnel is roughly estimated. DRI = 40.

76

5. BACK-MAPPING

5.3 Rock Sampling

5.3 ROCK SAMPLING The most important quality of a rock sample for laboratory testing, is that it is representative of the rock type. The selection of sample sites must therefore be carefully evaluated. The following procedures may be utilised to find the best sample sites: • Suitable sample sites with regard to representative and constant rock conditions as well as sampling possibilities (block or core sample) are recorded during the backmapping. • The number of samples is decided from the back-mapping and the machine performance data. Each rock type should be represented by at least one sample. One should at least test one rock sample for each 500 m of tunnel. • The sample site should be located in a 20 - 30 m long subsection of stable degree of fracturing and machine performance. • The sample site should be located where the Ic/In ratio is close to 1.0, see Table 2.3. • The sample site should not be close to larger discontinuities or intrusions, where the rock properties may be influenced by alteration. • Avoid sample sites with a high stress concentration along the tunnel circumference. The sampling method may vary, but block or core samples are recommended. Cores should have a diameter of 50 mm or more. One sample may consist of several pieces totalling 10 kg. Each piece should have a weight of at least 0.5 kg. Samples consisting of the largest chips from the boring may be used for laboratory testing, with the restrictions given in Section 4.6. It is a good practise to collect one or two large chips routinely (e.g. each day) to record the rock type, grain size, colour, etc. Typical chips from each 50 m tunnel section, each week or similar lengths or time periods are stored until the tunnel is finished, building a rock database for the tunnel.

77

5. BACK-MAPPING

5.4 Achieved vs. Predicted Performance

5.4 ACHIEVED VS. PREDICTED PERFORMANCE When the DRI, CLI, quartz content, ks-avg, kesp and the applied machine parameters have been found for a tunnel section, the achieved penetration rate, advance rate and cutter consumption may be compared to the prediction model in the Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear. Thus, the actual tunnelling performance may be evaluated against a standardised model. New data as a basis for improvement of the prediction models will also be provided in such cases. The model for penetration addition due to Marked Single Joints is based on less data than the penetration rate model for systematic fractured rock mass. Hence, when evaluating the machine performance as net penetration rate, one will get the best picture when using data from tunnel sections with few or no Marked Single Joints. The data from the back-mapping should also be compared to the predicted geological model established through the site investigations. Such verification will bring valuable information to improve: • The site investigation routines • The interpretations of surface mapping and core hole logging • The contract specifications and how to handle changed rock conditions.

78

APPENDIX

A. Previous Editions

A. PREVIOUS EDITIONS Previous editions of the Hard Rock Tunnel Boring Report including project group members: 1-76 Norwegian edition Bengt Drageset Roy-Egil Hovde Erik Dahl Johansen Roar Sandnes O. Torgeir Blindheim Odd Johannessen

1-79 Norwegian edition Knut Gakkestad Jan Helgebostad Svein Paulsen Oddbjørn Aasen Erik Dahl Johansen O. Torgeir Blindheim Odd Johannessen

1-83 Norwegian and English edition Arne Lislerud Steinar Johannessen Amund Bruland Tore Movinkel Odd Johannessen

1-88 Norwegian and English edition Arne Lislerud Amund Bruland Bjørn-Erik Johannessen Tore Movinkel Karsten Myrvold Odd Johannessen

1-94 Norwegian and English edition Bård Sandberg Amund Bruland Jan Lima Odd Johannessen

79

APPENDIX

B. Research Partners

B. RESEARCH PARTNERS The following external research partners have supported the project: • • • • • • • • • • •

Statkraft anlegg as Norwegian Public Roads Administration Statsbygg Scandinavian Rock Group AS NCC Eeg-Henriksen Anlegg AS Veidekke ASA Andersen Mek. Verksted AS DYNO Nobel Atlas Copco Rock Drills AB Tamrock OY The Research Council of Norway

80

APPENDIX


C. List of Parameters The parameters used in the report are listed in the following. The list is according to when the parameter first is explained or treated. Parameter Description af AV AVS b cc CAI CLI dc ds dtbm dtc DRI fch fD fD0 fhw fkd fwl hc,i hch Hf Hh hi Hm Hni

Unit

Page

Average spacing between planes of weakness cm Abrasion Value (tungsten carbide) mg/5 min (see also PR 13A-98) Abrasion Value Steel (cutter ring steel) mg/1 min (see also PR 13A-98) Penetration coefficient Cutter constant - cutterhead torque CERCHAR Abrasivity Index Cutter Life Index (see also PR 13A-98) Cutter diameter mm Sieve opening mm TBM or cutterhead diameter m Effective diameter of a thrust cylinder mm Drilling Rate Index (see also PR 13A-98) Chipping frequency Cutterhead factor - cutter wear Cutterhead factor of the reference cutterhead (1.133) Chip shape factor - thickness to width Kerf depth factor in chip forming Chip shape factor - width to length Machine hours of cutter change no. i h Individual and average largest chip thickness mm Cutter ring life - for the cutterhead sm3/c Cutter ring life - for the cutterhead h/c Measured intermediate ring height of an individual cutter ring at the cutterhead mm Cutter ring life - for the cutterhead m/c Number of cutter rings used at cutter position no. i

67 62 62 44 48 63 62 25 58 47 44 62 56 25 25 57 59 57 32 54 31 28 34 28 25

81

APPENDIX


Parameter Description

Unit

Page

hr Hri hwi

mm

34 25

mm mm mm A A A

35 35 34 36 14 14 14

m/h mm m/h m/h A m/h MPa

32 59 10 9 47 10 63

mm A m/week m/week mm/rev m/h

41 42 17 21 44 75 48

hwr h0 HRC IB IBj IBm Ic ik Im In IN Inj Is it It Iu Iun i0 I0 kc kD kesp ks ks-avg ks-i lc,i lch lesp

Measured ring height of replaced cutter ring Relative cutter ring life of cutter position no. i Intermediate wear height of an individual cutter ring at the cutterhead Wear height of replaced cutter ring Measured ring height of new cutter ring Rockwell C hardness number Applied amperage Amperage over subsection j Average amperage over a given tunnel length Net penetration rate - calculated from the cutter consumption Kerf depth at chip forming Average net penetration rate Net penetration rate Rated amperage of a cutterhead drive motor Net penetration rate of subsection j Point load strength index Measured penetration of one level in a penetration test Applied amperage of a penetration test level Weekly advance rate Normalised weekly advance rate Basic penetration Basic net penetration rate Cutter coefficient - cutterhead torque Correction factor for TBM diameter - cutter ring life Correction factor for Marked Single Joints Fracturing factor Average fracturing factor Fracturing factor of Fracturing Class i of a section Chainage of cutter change no. i mm Individual and average largest chip length Cumulated length of MSJs in an MSJ group of a section

m mm

25 74 72 73 73 32 54

m

75 82

APPENDIX


Parameter Description li lj Lt l1 l2 MB MBm MBj Mt M1 MSJ Nc nesp Ni nm Ntbm ntc nu pB pt ri rmc rri RPM S20 SJ Tb TB

Tunnel length of Fracturing Class i of a section Tunnel length of subsection j Total tunnel length Chainage at the start of a time period or tunnel section Chainage at the end of a time period or tunnel section Gross average thrust Gross average thrust over a given tunnel length Gross average thrust over subsection j Gross average thrust of a penetration test level Critical thrust to achieve a penetration rate of 1 mm/rev Abbreviation of Marked Single Joint Cutter change no. Number of MSJs in the MSJ group of a section Cutter position no. i Number of cutterhead drive motors Number of cutters on the cutterhead Number of thrust cylinders of the TBM Number of productive weeks of a period Applied thrust cylinder pressure Thrust cylinder pressure of a penetration test level Radius to cutter position no. i Relative radius to the position of the average cutter position Relative radius of cutter position no. i Cutterhead revolutions per minute Brittleness Value after 20 impacts (see also PR 13A-98) Sievers' J-value by miniature drill (see also PR 13A-98) Net time for boring Applied torque

Unit

Page

m m km

73 10 11

m

20

m kN/cutter kN/cutter kN/cutter kN/cutter

20 12 12 12 40

kN/cutter

week bar or psi

44 74 28 75 25 14 25 44 20 41

bar or psi m

41 48

rev/min %

48 25 14 62

mm/10

62

h kNm

16 14

83

APPENDIX


Parameter Description Tbak Tbj Tbr Tc Teh Tex Tmt Toa Tot Tp Tr Trs Ts Tsh tt Ttbm Tw t1 t2 ua UB um UN un Ut Vch wch

Unit

Time used for repair and service of the backup equipment h Machine hours used to bore subsection j h Time used for boring and regripping h Time used for cutter change and inspection h Effective working hours h/week (see also PR1B-98) Available time for tunnel excavation - excluding rock support work h Stop time due to rock transport problems h Stop time due to other activities h Stop time due to other transport problems h Stop time related to the tunnelling crew, such as travel, change of crews, lunch breaks, etc. h Time for regripping h Stop time due to rock support, water inflows and other geological causes h Stop time due to surveying h Total shift hours - working hours h Duration of one thrust level of a penetration test min Time used for repair and service of the TBM h Stop time due to water, electricity or ventilation problems h Machine hours at the start of a time period or tunnel section h Machine hours at the end of a time period or tunnel section h Achieved machine utilisation % Applied voltage of the cutterhead drive motors V Averaged machine utilisation % Rated voltage of the cutterhead drive motors V Normalised machine utilisation - per week of 100 shift hours % Applied voltage of the cutterhead drive motors during a penetration test V Cubic chip size mm3 Individual and average largest chip width mm

Page

15 10 15 15 17

17 15 15 15 15 16 15 15 16 41 15 15 16 16 17 14 18 42 17 42 56 54 84

APPENDIX


Parameter Description wh wj,i wj,i wm Ws cosφ⋅η α

αesp αf αs αt

Cutter wear rate - for the cutterhead Instantaneous cutter wear rate of cutter position no. i from Nc = j to Nc = j+1 Instantaneous cutter wear rate of cutter position no. i from Nc = j to Nc = j+1 Cutter wear rate - for the cutterhead Specific energy of rock cutting with TBM Efficiency factor of the cutterhead motors Angle between the tunnel axis and the planes of weakness Angle between the tunnel axis and Marked Single Joints Dip angle of a fracture or fracturing system Strike angle of a fracture or fracturing system Strike angle of the tunnel axis

Unit

Page

c/h

28

c/h

28

c/m c/m MJ/m3

28 28 50 14

°

72

° or ° g or ° g or °

74 66 66 66

g

85

APPENDIX

D1. Mapping Sheet ±135°

C h a in a g e

T U N N E L :

-1 3 5 °

-9 0 °

-4 5 °

0 °

4 5 °

9 0 °

D a te :

S ig n a tu r e :

1 3 5 °

D1. MAPPING SHEET ±135°°

L e ft w a ll

R o o f

R ig h t w a ll

R o c k ty p e

F r a c tu r in g

C o m m e n ts

The sheet is also available in various digital formats.

86

APPENDIX

D2. Mapping Sheet ±180°

C h a in a g e

T U N N E L :

-1 8 0 °

-9 0 °

-6 0 °

0 °

9 0 ° 6 0 °

D a te :

S ig n a tu r e :

1 8 0 °

D2. MAPPING SHEET ±180°°

L e ft w a ll

R o o f

R ig h t w a ll

R o c k ty p e

F r a c tu r in g

C o m m e n ts


87

2 0

C o m m e n ts :

B o r in g in c l. r e g r ip C u tte r c h a n g e a n d in s p e c tio n W a te r , e le c tr ic ity , v e n tila tio n T B M , r e p a ir a n d s e r v ic e B a c k - u p , r e p a ir a n d s e r v ic e C o n tin u o u s c o n v e y o r , r e p a ir a n d s e r v ic e A u x ilia r y tr a n s p o r t R o c k s u p p o rt W a te r in flo w S u r v e y in g L u n c h b re a k O th e r

T h ru s t

A m p e re

0 8

B o r e d le n g th :

B o r in g tim e :

1 9

C h a in a g e s ta r t:

M a c h in e h o u r s s ta r t:

0 7

C h a in a g e s to p :

M a c h in e h o u r s s to p :

D a y : N ig h t:

O p e ra to r:

T B M

D a te :

S H IF T L O G

2 1

0 9 2 2

1 0

6

4

2 1 1

0

2

4

6

8

2 3

2 4

1 2

6

4 6

4

2

2

0

0

2

4 2

6

6

S ta r t o f s h ift

4

6

L a s e r ta rg e ts : F ro n t

P R O J E C T :

0 1

1 3

4

2

0

B a c k

2 0 2

1 4

4

6

8 0 3

1 5

6

4

2

0

2

4

6

6

4 0 4

1 6

2

0

F ro n t

2

4 0 5

1 7

6

8

6

4

2

0

2

4

6

6 0 6

1 8

E n d o f s h ift

4

2

0

B a c k

0 7

1 9

2

4

6

8

APPENDIX D3. Shift Log

D3. SHIFT LOG


88

1 9

T o ta l n o . o f c u tte rs

M a c h in e h o u r s

C h a in a g e

1

6 .

2

1 1

7

4

1 2

2 5

12

1 7

D a te

3

6 .

1 1

2 4 2 2

9 .

C u tte r c h a n g e n o .

2 1

1 8

1 4

2 1

8

1 0

1 9

2 2 25

2 3

1 5

1 0

1 4

1 5

5

2 4

1 6

2 0

2 0

5

1 8

1 3

9 .

8

1 6

1 3

1 7 1 8 1 9 2 0 8

7

6

2 5

C o m m e n ts :

1 4

1 3

2 3

1 1

2 4

2 2

1 0 1 2

2 1 9

S H IF T :

1 6

P o s . n o .

5

R e a s o n

1 -3

C u tte r n o . o u t 1 5

in

1 -3

P o s . n o .

b in lo il ro th r in g s

e a r

e m a rk )

g

R e a s o n

r a s iv e ) w g c h ip p in c k e d b e a le a k a g e k e n b o lts e r ( w ith r

C u tte r n o . o u t

= (A = R = B = O = B = O

O P E R A T O R :

in

O

F

L

B

C

R E A S O N F O R C H A N G E : W

APPENDIX D4. Cutter Change Log

D4. CUTTER CHANGE LOG

2 3

1 7


89

P a s s in g %

0

0 .0 0 2

S ie v s iz e , m m

C L A Y

0 .0 0 0 6 0 .0 0 1

1 0

2 0

3 0

4 0

5 0

6 0

7 0

8 0

9 0

1 0 0

0 .0 0 5

0 .0 1

S IL T

0 .0 2

0 .0 6

0 .0 6 3 0 .1

0 .1 2 5 0 .2

0 .2 5

S A N D

0 .5

0 .5 1

1 2

2

4 5

8 1 0

1 6

G R A V E L

2 0

m m

3 2

4 5

6 0

6 4 1 0 0

APPENDIX D5. Sieve Curve Sheet

D5. SIEVE CURVE SHEET


90

APPENDIX

E. Penetration Test Spreadsheet

E. PENETRATION TEST SPREADSHEET A

B

C

D

E Normalised Penetration Curve

1

Mt

i0

log10(Mt)

log10(i0)

2

205.5

1.74

2.312811826

0.240549248

1.742153183

3

231.2

2.41

2.36398783

0.382017043

2.418340339

4

256.9

3.28

2.409764104

0.515873844

3.242800145

5

282.6

4.2

2.451172158

0.62324929

4.228296974

6

#NUM!

#NUM!

0

7

#NUM!

#NUM!

0

2.783161475

-6.195842434

8 9

ARX + BR

10 11

6.37027E-07

12 13

M1=

168.3405072

b=

2.783161475

12 15

Cell Entries A2:A7 B2:B7 C2:C7 D2:D7 B9 C9:D9

Applied thrust of two to six test levels Achieved basic penetration of two to six test levels Log values of the cutter thrust [=LOG10(A2)….=LOG10(A7)] Log values of the basic penetration [=LOG10(B2)….=LOG10(B7)] Text as shown {=LINEST(D2:D7,C2:C7,TRUE,FALSE)} Enter the formula as an array formula into both cells by: 1 Select the range C9:D9 2 Type the following formula: =LINEST(D2:D7,C2:C7,TRUE,FALSE) 3 Press Ctrl+Shift+Enter The formula will be displayed surrounded by brackets as shown above. D11 =10^D9 C13 Text as shown D13 =(1/D11)^(1/C9) C15 Text as shown D15 =C9 F2:F7 =(A2/D$13)^D$15…..=(A7/D$13)^D$15

91

ISBN 82-471-0281-1 ISSN 0802-3271