benchmarking final report

ICS-II, Task 4

GEOTECHNICAL GUIDELINES FOR A TRANSITION FROM OPEN PIT TO

UNDERGROUND MINING

Main Activity 1:

BENCHMARKING REPORT Prepared by

German Flores Antonio Karzulovic

December 2002

C O N T E N T S EXECUTIVE SUMMARY

1. INTRODUCTION 1

2. BENCHMARKING 2

3. DATA PROCESSING 4

4. GENERAL DATA 7

5. GEOTECHNICAL DATA 9

5.1. STRUCTURES 9

5.2. ROCK MASS 9

5.3. STRESS ENVIRONMENT 18

5.4. HYDROGEOLOGY 18

5.5. GEOTECHNICAL SOFTWARE 22

6. MINE DESIGN DATA 25

6.1. SLOPE GEOMETRY 25

6.2. MINE ACCESSES 25

6.3. BLOCK HEIGHT AND FOOTPRINT 30

6.4. CAVING INITIATION 32

6.5. UNDERCUT LEVEL 35

6.6. EXTRACTION LEVEL 35

6.7. SUPPORT 43

6.8. MATERIAL HANDLING SYSTEM 47

7. MINE OPERATION DATA 48

8. GEOTECHNICAL INSTRUMENTATION AND MONITORING DATA 51

9. GEOTECHNICAL HAZARDS DATA 53

9.1. COLLAPSES 53

9.2. ROCKBURSTS 56

9.3. SUBSIDENCE 61

9.4. WATER INFLOWS AND MUDRUSHES 68

9.5. HANGUPS 71

9.6. FINAL COMMENTS 73

10. CONCLUSIONS 76

11. ACKNOWLEDGMENTS 78

12. REFERENCES 79

Appendix A: GENERAL DATA ON MINES VISITED

Appendix B: BENCHMARKING SURVEYS

Appendix C: DATABASE

EXECUTIVE SUMMARY

In the next 10 to 15 years several mines are considering a transition from open pit to underground cave mining. These include: Argyle Diamond Mine, Bingham Canyon, Chuquicamata, Grasberg, WMC Mount Keith and Newcrest Telfer. Considering this fact, the ICS-II included Task 4 with the goal of providing the project sponsors with practical geotechnical guidelines to develop the transi-tion from open pit to underground cave mining.

To achieve this objective, the following main activities have been considered: Benchmarking, Geo-technical Guidelines, worked Example and Final Report. Currently, and according to the program approved at the ICS-II Meeting of October 2001, in Santiago, only the first activity, BENCHMARK-ING, has been developed and it is presented in this report.

The benchmarking study was planned and developed according to a program aimed to optimize data collection:

1. SURVEY DESIGN: This was the first task to be completed. In order to facilitate data collec-tion, an Excel spreadsheet was designed, and e-mailed to the targeted mines also willing to provide information.

2. MINE VISITING: 17 mines were selected to be visited and relevant information was ob-tained. The selection criterion was mines which have developed, or are planning to develop, a transition from open pit to underground mining, and also other mines (open pits and un-derground) that could provide relevant information.

3. ADDITIONAL DATA COLLECTION: a comprehensive survey of the available technical lit-erature was done in order to collect supplementary data. This allows the inclusion of data on 88 additional mines; nevertheless, in most of the cases, the additional data does not include all the features considered in the benchmarking survey.

4. DATA PROCESSING: The collected data was analyzed in order to develop histograms and, where possible, correlations showing the current practices and trends of underground mining by caving methods. When enough data was available the relative frequency of the different parameters was computed, and when the available data was limited, the relative importance of the different parameters was assessed.

5. BENCHMARKING REPORT: All of the above mentioned, and the conclusions and recommendations resulting from this benchmarking are presented in this report.

The interpretation of the data collected in this benchmarking has allowed to define the current trends and practices of the underground mining by caving methods. These have been summarized as histograms and/or curves to facilitate their use by the sponsors of ICS-II, especially during the early stages of a new mining project.

One of the main results of this study is shown in the following Table which summarizes the current trends for the most relevant design parameters used at the caving mine operations.

Finally it must be noted that all the results presented in this report will be used as a starting basis for the development of geotechnical guidelines for a transition from open pit to underground mining, which corresponds to the second main activity of Task 4, and includes the following subjects:

1. CAVING PROPAGATION

2. SUBSIDENCE

3. CROWN-PILLAR

4. WATER INFLOWS

TYPICAL DESIGN PARAMETERS FOR A BLOCK/PANEL CAVING MINE

Mine Design Parameter Typical Value Comments

Rock Mass Quality 50 RMR < 60 If RMR > 60 rock mass cavability must be evaluated carefully.

Acces Decline Currently 70% of mines prefer declines, and 20% declines and shafts as mine access.

Block Height 210 m This typical block height could vary 20%.

< 50000 m2 30000 m2 50000 a 100000 m2 75000 m2 Footprint Area

> 100000 m2 170000 m2

These typical areas could vary +20%. It is recommended to use equal or larger areas, but not smaller than the typical values. Also, square areas are better than the rectangular ones.

Area 10000 m2 Smaller areas are not recommended, specially in massive rock masses.

Shape Square Internal corners must be avoided (e.g. a L shaped area).

Measures to Facilitate Slot Is highly recommended to facilitate cave initiation. Caving

Initiation

Hydraulic Radius 20 to 30 m Avoid being close to the limit in Laubschers chart.

Spacing 15 m This is the current practice. Height 4 m D

rifts

Width 4 m Could be increased but not decreased.

Undercut Height 8 m Could vary, but be careful if using small undercutting heights.

Und

ercu

t Le

vel

Undercut Rate 2100 m2/month Could be increased but be careful with induced seismicity, specially if in a high stress environment.

Crown-Pillar Thickness 17 m Could vary 20% (measured from floor UCL to floor EXT).

Spacing 30 m Could vary from 26 to 36 m. Height 4 m Ex

tract

ion

Leve

l

Drif

ts

Width 4 m Could be increased but not decreased.

Spacing 15 m Could vary from 13 to 18 m. Draw Points

Influence Area 225 m2 Could vary from 169 to 324 m2.

Draw Rates 0.20 m/day This is an average value. Typically lower values are used at the beginning of caving, and higher values are used when over 30% of the block height has been extracted.

Capacity 11 ton It could vary 20%. LHD Equipment Traming Distance 140 m Smaller tramming distances are preferable.

Powder Factor 400 grm/ton For undercutting blasting. It could vary 20%.

Oversize Limit 1.8 to 2.0 m3 It could vary 20%. RMR < 70 > 45 Subsidence RMR > 70 > 60 is the break angle defining the mean inclination of the crater walls.

Geotechnical Hazards The project must take account that collapses, rockbursts, subsidence, water inflows and mudrushes, and hangups could occur

Instrumentation & Monitoring The most common monitoring systems include displacements and seismicity. It is re-commended to include a seismic monitoring system, specially in massive hard rock and/or high stress environments..

(1) These typical values are intended only for the pre-feasibility stage of a mining project. (2) RMR values are for Laubschers 1990 system.

ICS-II, Task 4 GEOTECHNICAL GUIDELINES FOR A TRANSITION FROM OPEN PIT TO UNDERGROUND MINING

1

1. INTRODUCTION

Several mines are considering a transition from open pit to underground cave mining, in the mid to long term. These include: Argyle Diamond Mine, Bingham Canyon, Chuquicamata, Grasberg, WMC Mount Keith and Newcrest Telfer. Considering this fact, the ICS-II included Task 4 with the goal of providing the project sponsors with practical geotechnical guidelines to develop the transi-tion from open pit to underground cave mining.

To achieve this objective, the following has been considered:

1. BENCHMARKING, to collect data from mines which have developed, or are planning to de-velop, a transition from open pit to underground mining, and also from other mines (open pits or underground) that could provide relevant information for this research. The collected data was supplemented by a comprehensive review of the available technical literature.

2. GEOTECHNICAL GUIDELINES, to develop practical methodologies to deal with the key is-sues arising in a transition from open pit to underground cave mining. These guidelines will address the following subjects: Caving Propagation, Subsidence, Crown/Buffer-Pillar, and Water Inflows.

3. WORKED EXAMPLE, to illustrate the use of these geotechnical guidelines by applying them to a real case example: Chuquicamata Mine.

4. FINAL REPORT, to include the results of the benchmarking, the geotechnical guidelines, and the worked example in a self-contained technical report.

Currently, and according to the program approved at the ICS-II Meeting of October 2001, in Santi-ago, only the first activity, BENCHMARKING, has been developed and it is presented in this report.


2

2. BENCHMARKING

The benchmarking study was planned and developed according to a program aimed to optimize data collection:

6. SURVEY DESIGN: This was the first task to be completed. In order to facilitate data collec-tion, an Excel spreadsheet was designed, and e-mailed to the targeted mines also willing to provide information. These spreadsheets are included, with the data collected, in Appen-dix B.

7. MINE VISITING: 17 mines were selected to be visited and relevant information was ob-tained. The selection criterion was mines which have developed, or are planning to develop, a transition from open pit to underground mining, and also other mines (open pits and un-derground) that could provide relevant information. Table 2.1 summarizes the mines that were visited, and in Appendix A general information on these mines is presented.

8. ADDITIONAL DATA COLLECTION: a comprehensive survey of the available technical lit-erature was done in order to collect supplementary data. This allows the inclusion of data on 88 additional mines; nevertheless, in most of the cases, the additional data does not include all the features considered in the benchmarking survey.

9. DATA PROCESSING: The collected data was analyzed in order to develop histograms and, where possible, correlations showing the current practices and trends of underground mining by caving methods. When enough data was available the relative frequency of the different parameters was computed, and when the available data was limited, the relative importance of the different parameters was assessed. The databases resulting from this data process-ing are included in Appendix C.

10. BENCHMARKING REPORT: All of the above mentioned, and the conclusions and recom-mendations resulting from this benchmarking are presented in this report.


3

Table 2.1 MINES VISITED FOR BENCHMARKING

Country Mine Comments

Cadia Hill Open pit mine. Visited October 2002.

Mount Keith Open pit mine.

Project for a transition to underground mining. Visited May 2002

Northparkes Mine that developed a transition from open pit to underground mining.

Underground mining by block caving. Visited October 2002

Australia

Ridgeway Underground mining by sublevel caving. Visited October 2002

Canada Kidd Creek Mine that developed a transition from open pit to underground mining.

Underground mining by open stoping. Visited June 2002

Andina Open pit mine and underground mining by panel caving. Visited July 2002

Chuquicamata Open pit mine.

Project for a transition to underground mining. Visited July 2002

El Teniente Underground mining by panel caving. Visited July 2002

Chile

Salvador Underground mining by panel caving. Visited July 2002

Grasberg Underground (DOZ)

Indonesia Grasberg Open Pit

Underground mining by panel caving. Open pit mine.

Project for a transition to underground mining. Visited April 2002

Finsch Mine that developed a transition from open pit to underground mining.

Underground mining by open stoping. Visited May 2002

Koffiefontein Mine that developed a transition from open pit to underground mining.

Underground mining by sublevel / front caving. Visited May 2002

South Africa

Palabora Mine developing a transition from open pit to underground mining.

Open pit mine and underground mining by panel caving. Visited May 2002

Sweden Kiruna Mine that developed a transition from open pit to underground mining.

Underground mining by sublevel caving. Visited June 2002

Bingham Canyon Open pit mine.

Project for a transition to underground mining. Visited June 2002 USA

Henderson Underground mining by panel caving. Visited June 2002


4

3. DATA PROCESSING

The collected data was analyzed in order to develop histograms and, if possible, correlations show-ing the current practice and trends of open pit and underground mining by caving methods. When enough data was available the relative frequency of the different parameters was computed, and when the available data was limited the relative importance of the different parameters was as-sessed.

The collected data included: GENERAL DATA GEOTECHNICAL DATA MINE DESIGN DATA MINE OPERATION DATA MONITORING DATA GEOTECHNICAL HAZARDS DATA

The process of data collection and processing showed that the number of mines that have devel-oped, are in the process of developing, or will develop a transition from open pit to underground mining, or vice versa, was more than what was expected. Indeed, Table 3.1 summarizes data on 33 mines that are under this condition.

Also, the analysis of the data indicated a sudden increase in the pit depths of the mines that will have this transition in a mid or a long term, as illustrated by Figure 3.1. This is especially important because it means that the geotechnical challenges for these projects will be expected to be larger than the ones of the mines that had developed a transition in the past. The pits that will have large depths when initiating the transition process are:

Bingham Canyon, USA (747 to 849 m depth)

Chuquicamata, Chile (1100 m depth)

Grasberg, Indonesia (1000 m depth)


5

Table 3.1 MINES THAT DEVELOPED, ARE DEVELOPING OR WILL DEVELOP

A TRANSITION FROM OPEN PIT TO UNDERGROUND MINING

H PIT Date Country Mine Transition Type (m) Developments Mining

Argyle Diamond Open pit UG mining 150 a 300 2006 (?) Big Bell UG mining Open pit UG mining 1994 1997

Mount Isa 146 1967

Mount Keith 344 2015 (?)

Northparkes 100 1993 1997

Australia

Perseverance

Open pit UG mining

Craigmont 76 1963 1964

Kidd Creek 250 1969 1973

Stobie 150 (?) 1941 1948 Canada

Williams

Open pit UG mining

Chuquicamata 1100 2016 (?) Chile

Mansa Mina Open pit UG mining

400 (?) 2014 (?)

Finland Pyhasalmi Open pit UG mining 135 1967 Indonesia Grasberg Open pit UG mining 1000 2016 (?)

Kirovsky 1959 Russia

Mir Open pit UG mining

455 1994

Finsch 423 1979 1990

Koffiefontein 240 1981

Palabora 803 1996 2000

Premier 189 1945 1946 (?)

Thabazimbi 70 a 240 1988 (?)

South Africa

Venetia

Open pit UG mining

360 (?) 2011 (?)

Sweden Kiruna Open pit UG mining 230 1958 Bingham Canyon Open pit UG mining 747 a 899 2012 (?)

Climax 1973

Miami UG mining Open pit

San Manuel UG mining Open pit UG mining USA

Questa Open pit UG mining 150 (?) 1979 1983 Zambia Nchanga UG mining Open pit UG mining 1937 1939

Gaths

Miriam 60 1955 1957

Shabanie 150 (?) 1950 (?) Zimbabwe

Shangani

Open pit UG mining

150 1980 (?)


6

1945 1950 1955 1960 1965 1970 1975 1980 1985 1990 1995 2000 2005 2010 2015 2020

YEAR

0

100

200

300

400

500

600

700

800

900

1000

1100

1200

MA

XIM

UM

PIT

DEP

TH

(m)

CHUQUICAMATA

BINGHAM CANYON

GRASBERG

PALABORA

MINES THAT DEVELOPED A TRANSITION FROM OP TO UG MINES THAT ARE DEVELOPING A TRANSITION FROM OP TO UG MINES THAT WILL DEVELOP A TRANSITION FROM OP TO UG TREND FROM CASE HISTORIES TREND FROM PROJECT DATA

Figure 3.1: Evolution through time of the trend for the depth of open pit mines that have devel-oped, are developing, or will develop a transition to underground mining.


7

4. GENERAL DATA

The general data collected include: Mine name, location, country, and owner. Mine elevation, ore type, mined out reserves, initial mining method, and initial mining date. Current mining method, reserves, mean ore grade, mine life, final mine depth, cash and total

costs (if provided), ore production, and waste removal. Future mining method, reserves, mean ore grade, mine life, final mine depth, cash and total

costs (if provided). Total work force. Geotechnical groups (engineers, geologists, technicians) Additional comments. All the data obtained for each mine visited are included in Appendix B.

The analysis of the production and information on geotechnical groups is summarized in Figure 4.1, and indicates that:

(a) Due to the nature of mining methods open pit mines have much larger ore production than underground mines; therefore, any open pit considering a transition to underground mining must take account of this fact.

(b) Typically geotechnical groups are larger in underground mines than in open pit mines (of course there are a few exceptions).

(c) According to the data, it is possible to define a trend between the size of the typical geotech-nical group and the ore production for open pit and underground mining. These trends indi-cate that:

The larger the ore production the larger the typical geotechnical group in both cases, open pit and underground mining.

This trend shows a break or a sudden increase in the number of people in the geo-technical group when the ore production exceeds 25 kTPD in underground mines, and 75 kTPD, in open pit mines.

(d) Therefore, considering a transition to underground mining, any open pit must take this fact into account, and probably will have to increase the number of people in its geotechnical group (in spite of the fact that the underground ore production will be smaller that the open pit production).


8

0 25 50 75 100 125 150 175 200 225 250

ORE PRODUCTION (kTPD)

0

5

10

15

20

25

30

GEO

TEC

HN

ICA

L G

RO

UP

(pe

ople

)

MINING METHOD

OPEN PIT UNDERGROUNG

MASS

IVE O

PEN

PIT M

INING

(more

than

75 kT

PD)

MAS

SIVE

UND

ERG

ROUN

D M

ININ

G

(mor

e th

an 2

5 kT

PD)

OP (< 75 kTPD)UG (< 25 kTPD)

Figure 4.1: Variation of the size of the typical geotechnical group with ore production in open pit

and underground mining.


9

5. GEOTECHNICAL DATA

The geotechnical data collected included information on Structures, Rock Mass, Stress Environ-ment, and Hydrogeology.

5.1. STRUCTURES

The data collected on structures include: Structural domains Number of structural sets Geological characteristics: type of structure, infilling, waviness, roughness, and water

condition. Geometrical characteristics: dip, dip direction, length, spacing, and gap. Mechanical properties: joint roughness coefficient (JRC), joint wall compressive

strength (JCS), dilation angle (i), cohesion (cJ), friction angle (J), normal stiffness (kN), and shear stiffness (kS).

All the data obtained for each mine visited are included in Appendix B.

The analysis of the data on the orientation and properties of the structural sets in open pit and underground mines indicates that:

(a) In most underground mines that use caving methods, subvertical structures predomi-nate (subvertical meaning dips steeper than 60), as shown in Figures 5.1 and 5.2. This conclusion does not mean that there are not subhorizontal or flatter structures, but that the number of subvertical sets (> 60) exceeds the number of flatter sets (< 60).

(b) In underground and open pit mines the data on the orientation of structures is typically much better than the data on their length, spacing, and gap. Generally the data can be ordered from more to less reliable as follows:

Dip Dip Direction Spacing Length Gap

(c) The geotechnical characterization of structures generally is poorer in underground mines than in open pit mines. Perhaps due to the fact that mapping is more difficult underground. This is shown in Figure 5.3 that correlates the magnitude of the cohe-sion and friction angle, and shows a much better trend in the data from open pits than in the one from underground mines.

(d) In open pit mines the strength properties of structures are fairly to well known, but the deformability properties are poorly to fairly known.

(e) In underground mines the strength properties of structures are poorly to fairly known, but their deformability properties are almost unknown.

(f) In spite of the increasing use of numerical models, the quality of input data on the me-chanical properties of structures is, in most of cases, poor.

5.2. ROCK MASS

The data collected on rock masses include: Rock types. Intact rock properties: unit weight (), uniaxial compressive strength (UCS), parameter

m of the Hoek-Brown criteria (mi), modulus of deformability (E), wave velocity for P and S waves (VP and VS).


10

85 - 90 1 1 3 2 1 1 3 1 2 1 3

80 - 84 2 4 3 1 1 1 3 4 4 3 2 1 2

75 - 79 1 1 2 1 2 1 1 3 1 2 2 1 1 2

70 - 74 1 1 4 1 1 2 1 1 1 4 2 1

65 - 69 2 1 1 6 1 2 11 3 1 3 3

60 - 64 1 1 2 1 1

55 - 59 1 1 1 1 3 1 2

50 - 54 1 1 1 1 1 1

45 - 49 2 1 1 1 2 1 1

40 - 44 1 1

35 - 39 1 3 1

30 - 34 1

25 - 29

20 - 24 1

15 - 19 1

10 - 14 1

5 - 9 1

0 - 4

0 - 19 20 - 39 40 - 59 60 - 79 80 - 99 100 - 119 120 - 139 140 - 159 160 - 179 180 - 199 200 - 219 220 - 239 240 - 259 260 - 279 280 - 299 300 - 319 320 - 339 340 - 359

DIP

D I P D I R E C T I O N Figure 5.1: Trend of the orientation (defined by dip and dip direction) of structural sets in under-

ground mines that use caving methods.

0.00 0.02 0.04 0.06 0.08 0.10 0.12 0.14 0.16 0.18 0.20 0.22

RELATIVE FREQUENCY

90

80

70

60

50

40

30

20

10

0

DIP

(d

egre

ss)

Figure 5.2: Histogram showing the relative frequency of different dip angles for structural sets in

underground mines that use caving methods.


11

Rock mass quality: RQD, RMRBIENIAWSKI, RMRLAUBSCHER, Q, GSI. Rock mass properties: cohesion (c), friction angle (), modulus of deformability (E),

Poissons ratio (), bulk modulus (B), shear modulus (G), wave velocity for P and S waves (VP and VS).


The analysis of the data on the rock masses in open pit and underground mines indicates that:

(a) The data on intact rock properties is well known for the unit weight (), and the uniaxial strength (UCS); but the data for the other intact rock parameters is poorer.

(b) Typically UCS values are smaller for open pit mines rocks (averages 80 MPa) that for underground mines rocks (averages 115 to 150 MPa). There is also no major differ-ence in the UCS values for the rocks in different types of underground mining. This is shown in Figure 5.4.

(c) Typically RQD values are smaller for open pit mines rocks (averages 65%) than for underground mines rocks (averages 70% to 85%). Also there is no major difference in the RQD values for the rocks in different types of underground mining. This is shown in Figure 5.5.

(d) The most used method for rock mass classification in underground mines is Laub-schers RMR (53%), followed by Bartons Q (26%), and Bieniawskis RMR (15%). The most used method for rock mass classification in open pit mines is Hoeks GSI (39%), followed by Bieniawskis RMR (26%), and Laubschers RMR (22%). This is shown in Figure 5.6.

15 20 25 30 35 40 45

FRICTION ANGLE OF GEOLOGICAL STRUCTURES (degrees)

0

25

50

75

100

125

150

175

200

CO

HES

ION

OF

GEO

LOG

ICA

L S

TRU

CTU

RES

(k

Pa) OPEN PIT MINING

UNDERGROUND MINING

Figure 5.3: Variation of the cohesion of structures with their friction angle, for open pit and un-

derground mines.


12

0 50 100 150 200 250 300 350 400 450 500

INTACT ROCK UNIAXIAL COMPRESSIVE STRENGTH, UCS (MPa)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

RE

LATI

VE

FR

EQ

UEN

CY

MINING METHOD

OPEN PIT OPEN STOPING SUBLEVEL CAVING BLOCK CAVING PANEL CAVING

Figure 5.4: Relative frequency of the intact rocks uniaxial compressive strength, UCS, in differ-

ent mining methods.

0 10 20 30 40 50 60 70 80 90 100

ROCK QUALITY DESIGNATION, RQD (%)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

RE

LATI

VE

FR

EQ

UEN

CY

MINING METHOD


Figure 5.5: Relative frequency of the Rock Quality Designation Index, RQD, in different mining

methods.


13

(e) Interpreting all rock mass classification data in terms of Laubschers RMR it is clearly

evident, as shown in Figure 5.7, that rock mass quality is poorer in open pit mines (averages 40) than in underground mines (averages 50 to 60).

(f) The typical rock mass rating distribution for different mining conditions are shown in Figures 5.8 to 5.12, which show the following typical RMR ranges:

Open Pit Mines RMR: 20 to 40 Open Stoping Mines: RMR: 40 to 80 Sublevel Caving Mines: RMR: 40 to 70 Block Caving Mines: RMR: 30 to 70 Panel Caving Mines: RMR: 40 to 80

(g) As shown in Figure 5.13, the average trend relating Laubschers RMR and MRMR is:

MRMR = 0.9 RMR

(h) As shown in Figure 5.14 the cohesion of underground mines rock masses is typically larger than the cohesion of open pit rock masses, probably due to the higher confine-ment in underground mining. This figure also shows that the trend between rock mass cohesion and rock mass friction angle is better for the case of open pits than for un-derground mines.

(i) The geotechnical characterization of rock masses seems to be poorer in underground mining than in open pit mining. Indeed, in spite of the increasing use of numerical models the quality of input data on rock mass properties is, in most cases, poor to fair.

ROCK MASS CLASSIFICATION SYSTEM

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

RE

LATI

VE F

REQ

UE

NC

Y

MINING METHOD

OPEN PIT UNDERGROUND

Q (Barton et al.) RMR (Bieniawski) GSI (Hoek et al.) RMR (Laubscher)

Figure 5.6: Methods used in mining for rock mass classification.


14

0 10 20 30 40 50 60 70 80 90 100

LAUBSCHERS ROCK MASS RATING, RMR

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

RE

LATI

VE

FR

EQ

UEN

CY

MINING METHOD


Figure 5.7: Relative frequency of Laubschers Rock Mass Rating, RMR, in open pits and under-

ground mines that use different mining methods.

0 10 20 30 40 50 60 70 80 90 100


0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

RE

LATI

VE

FR

EQ

UEN

CY

Figure 5.8: Relative frequency of Laubschers Rock Mass Rating, RMR, in open pit mining.


15

0 10 20 30 40 50 60 70 80 90 100


0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

RE

LATI

VE

FR

EQ

UEN

CY

Figure 5.9: Relative frequency of Laubschers Rock Mass Rating, RMR, in open stoping mining.

0 10 20 30 40 50 60 70 80 90 100


0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

RE

LATI

VE

FR

EQ

UEN

CY

Figure 5.10: Relative frequency of Laubschers Rock Mass Rating, RMR, in sublevel caving mining.


16

0 10 20 30 40 50 60 70 80 90 100


0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

RE

LATI

VE

FR

EQ

UEN

CY

Figure 5.11: Relative frequency of Laubschers Rock Mass Rating, RMR, in block caving mining.

0 10 20 30 40 50 60 70 80 90 100


0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

RE

LATI

VE

FR

EQ

UEN

CY

Figure 5.12: Relative frequency of Laubschers Rock Mass Rating, RMR, in panel caving mining.


17

0 10 20 30 40 50 60 70 80 90 100

LAUBSCHERS RMR

0

10

20

30

40

50

60

70

80

90

MR

MR

MRMR / RMR = 1.2 1.1 1.0 0.9

MINING METHOD


0.8

0.5

0.7

0.6

Figure 5.13: Relationship between Laubschers Rock Mass and Mining Rock Mass Ratings, RMR

and MRMR.

15 20 25 30 35 40 45 50 55 60 65 70

ROCK MASS FRICTION ANGLE (degrees)

10

100

1000

10000

RO

CK

MA

SS

CO

HES

ION

(k

Pa)

OPEN PIT MINING UNDERGROUND MINING

2000

5000

500

50

200

20

Figure 5.14: Relationship between the cohesion and the friction angle of the rock mass in open pit

and underground mining.


18

5.3. STRESS ENVIRONMENT

The data collected on the stress environment include: Production sector where the stress measurements were made. Stress tensor components: horizontal stress (Sx, Sy, X towards East, Y towards North),

vertical stress (Sv), and shear stresses (Sxy, Syz, Szx). Principal stresses: magnitudes (S1, S2, S3), plunges (1, 2, 3), and trends (1, 2, 3). Stress measurement method. All the data obtained for each mine visited are included in Appendix B.

The analysis of the data on the stress environment in underground mines indicates that:

(a) Currently the CSIRO Hollow Inclusion Cell is the most used method for in situ stress measurements.

(b) As shown in Figure 5.15, in underground mines the in situ major principal stress S1 typically varies from 30 to 40 MPa.

(c) As shown in Figure 5.16, the minimum principal stress S3 typically varies from 10 to 20 MPa.

(d) As shown in Figure 5.17, the principal stress difference S1 - S3 typically varies from 20 to 30 MPa.

(e) As shown in Figure 5.18, in underground mines the in situ vertical stress is larger than the lithostatic stress (z). This result could be due to the fact that several stress measurements could be located in proximity to caves.

(f) As shown in Figure 5.19, in underground mines the mean value of the stress ratio, KMEAN, is bounded as proposed by Hoek & Brown (1980):

0.5 + (1500 / z) KMEAN 0.3 + (100 / z)

(g) As a result of this benchmarking, similar relationships were derived for the minimum and maximum values of the stress ratio, KMIN and KMAX. These relationships are shown in Figures 5.20 and 5.21, and are given by:

0.6 + (1250 / z) KMIN 0.2 + (100 / z)

1.0 + (1500 / z) KMAX 0.3 + (90 / z)

5.4. HYDROGEOLOGY

The data collected on the hydrogeology include: Hydrogeological units. Maximum and minimum permeabilities (kMAX and kMIN). General parameters: depth of the phreatic surface, infiltration rate into the mine, and

dewatering rate. Operative parameters on drainage systems: drainage tunnels, pumping wells, and

subhorizontal drains. All the data obtained for each mine visited are included in Appendix B.

The analysis of the data on the hydrogeology in open pits and underground mines indicates that:

(a) Most mines do not consider the hydrogeological characterization a high priority.


19

0 10 20 30 40 50 60 70 80 90 100

MAJOR PRINCIPAL STRESS, S1 (MPa)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

REL

ATI

VE F

RE

QU

ENC

Y

Figure 5.15: Histogram showing the relative frequency of major principal stresses, S1, with different

magnitudes (measurements in underground mines).

0 5 10 15 20 25 30 35 40

MINOR PRINCIPAL STRESS, S3 (MPa)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

REL

ATI

VE F

RE

QU

ENC

Y

Figure 5.16: Histogram showing the relative frequency of major principal stresses, S3, with different

magnitudes (measurements in underground mines).


20

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70

PRINCIPAL STRESS DIFFERENCE, S1 - S3 (MPa)

0.00

0.05

0.10

0.15

0.20

REL

ATI

VE F

RE

QU

ENC

Y

Figure 5.17: Histogram showing the relative frequency of major principal stress differences, S1 - S3,

with different magnitudes (measurements in underground mines).

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80 85 90

VERTICAL STRESS (MPa)

3000

2750

2500

2250

2000

1750

1500

1250

1000

750

500

250

0

DEP

TH

(met

ers)

Figure 5.18: Variation of in situ vertical stresses with depth in underground mines.


21

0.00 0.25 0.50 0.75 1.00 1.25 1.50 1.75 2.00 2.25 2.50 2.75 3.00

K MEAN

3000

2750

2500

2250

2000

1750

1500

1250

1000

750

500

250

0

DEP

TH

(met

ers)

Figure 5.19: Variation of the average value of the in situ stress ratio, KMEAN, with depth in under-

ground mines. The black curves shown the upper and lower boundaries defined by Hoek & Brown (1980), while the red curve is the average between them.

0.00 0.25 0.50 0.75 1.00 1.25 1.50 1.75 2.00 2.25 2.50 2.75 3.00

K MIN

3000

2750

2500

2250

2000

1750

1500

1250

1000

750

500

250

0

DEP

TH

(met

ers)

Figure 5.20: Variation of the minimum value of the in situ stress ratio, KMIN, with depth in under-

ground mines. The black curves shown the upper and lower boundaries defined in this work, while the red curve is the average between them.


22

(b) As shown in Figure 5.22, in open pit mines the most used drainage systems are:

subhorizontal drains (38%), drainage tunnels (27%), pumping wells (21%), and sumps (14%).

(c) As shown in Figure 5.22, in underground mines the most used drainage systems are: sumps (78%), subhorizontal drains (14%), and drainage tunnels (8%).

(d) The most typical monitoring systems are: observation wells (open holes), piezometers, and flow rate measurement devices.

5.5. GEOTECHNICAL SOFTWARE

The data collected on geotechnical software currently being used in open pit and under-ground mines, included the name and type of software. All the data obtained for each mine visited are included in Appendix B. The analysis of this data indicates that:

(a) As shown in Figure 5.23, for conventional slope stability analyses the most used soft-ware are: SLIDE (30%), DIPS (20%), and SWEDGE (17%).

(b) As shown in Figure 5.24, for two-dimensional numerical analyses the most used soft-ware are: FLAC (50%), UDEC (33%), and EXAMINE (10%).

(c) As shown in Figure 5.25, for three-dimensional numerical analyses the most used software are: FLAC3D (44%), 3DEC (26%), and MAP3D (18%).

0.00 0.25 0.50 0.75 1.00 1.25 1.50 1.75 2.00 2.25 2.50 2.75 3.00 3.25 3.50

K MAX

3000

2750

2500

2250

2000

1750

1500

1250

1000

750

500

250

0

DEP

TH

(met

ers)

Figure 5.21: Variation of the maximum value of the in situ stress ratio, KMAX, with depth in under-

ground mines. The black curves shown the upper and lower boundaries defined in this work, while the red curve is the average between them.


23

DEWATERING SYSTEM

0.0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

REL

ATI

VE F

REQ

UEN

CY

UNDERGROUND MINING OPEN PIT MINING

DRAINAGE TUNNELS PUMPING WELLS SUBHORIZONTAL DRAINS SUMPS

Figure 5.22: Relative frequency of different dewatering systems used in open pits and under-

ground mines.

SOFTWARE PACKAGE

0.00

0.05

0.10

0.15

0.20

0.25

0.30

RE

LATI

VE

FR

EQ

UEN

CY

SLIDE DIPS SWEDGE XSTABL ROCFALL BACKBREAK GALENA SLOPE/W UTEXAS NFOLD

CONVENTIONAL SLOPE STABILITYSOFTWARE CURRENTLY USED IN

OPEN PIT MINES

Figure 5.23: Relative frequency of software used in open pit mines for conventional slope stability

analyses.


24

SOFTWARE PACKAGE

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

RE

LATI

VE

FR

EQ

UEN

CY

FLAC UDEC EXAMINE / EXAMINE TAB PHASE2

2D NUMERICAL ANALYSISSOFTWARE CURRENTLY USED IN

OPEN PIT AND UNDERGROUND MINES

Figure 5.24: Relative frequency of software used in open pit and underground mines for two-dimensional numerical analyses.

SOFTWARE PACKAGE

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

RE

LATI

VE

FR

EQ

UEN

CY

FLAC3D 3DEC MAP3D EXAMINE3D BEFE ELAST-3

3D NUMERICAL ANALYSISSOFTWARE CURRENTLY USED IN

OPEN PIT AND UNDERGROUND MINES

Figure 5.25: Relative frequency of software used in open pit and underground mines for three-

dimensional numerical analyses.


25

6. MINE DESIGN DATA

On open pit mines the mine design data collected included information on Slopes Geometrical Pa-rameters, Acceptability Criteria, and Tools for Analysis.

On underground mines the mine design data collected included information on Mine Accesses, Mining Method, Cave Initiation, Footprint, Block Height, Mine Layout, and Materials Handling Sys-tems.


6.1. SLOPE GEOMETRY

The analysis of the data on slope geometries in open pit mines indicates that:

(a) As shown in Figure 6.1, bench heights can vary from 10 to 20 m for single benches, and from 25 to 35 m for double benches. The typical height for single benches is 15 m, while it varies from 25 to 30 m for double benches. In most cases double benches are developed in two stages (i.e. first a single bench is developed, and then it is dou-bled). This practice is very common for pushbacks that reach the final pit condition, and where the rock mass has a good geotechnical quality.

(b) As shown in Figure 6.2, in open pit slopes the interramp height can vary widely, from 50 to 250 m; but typically it does not exceed 200 m, and its average value is about 140 m.

(c) As shown in Figure 6.3, the overall height of open pit slopes can vary widely, from 100 to 900 m; but in most of the cases it varies from 100 to 500 m (more than 70% of the cases), and its average value is about 350 m.

(d) As shown in Figure 6.4, the bench face inclination can vary from 55 to 90; but in most of the cases it varies from 65 to 80, and its average is about 73. It is important to indicate that to achieve bench face inclinations steeper than 65, it is a common practice to use controlled blasting techniques.

(e) As shown in Figure 6.5, the interramp angle can vary from 25 to 60; but in most of the cases it varies from 40 to 60, and its average is about 50.

(f) As shown in Figure 6.6, the overall slope angle can vary from 25 to 60; but in most of the cases it varies from 30 to 60, and its average is about 45.

(g) As shown in Figure 6.7, the slope angle is maximum at bench scale, flatter for inter-ramp slopes (typically 20 to 25 flatter), and even flatter for overall slopes (typically 5 flatter than interramp slopes).

(h) As shown in Figure 6.8, the data for interramp and overall slopes do not show a clear trend between the slope height and the slope angle (probably due to the fact that the data include many different geological-structural-geotechnical settings); nevertheless, for preliminary evaluations the red curve shown in Figure 6.8 could be used to esti-mate the slope angle for a given slope height.

6.2. MINE ACCESSES

The analysis of the data on mine accesses indicates that:

(a) Underground mine accesses can be shafts, declines or both.

(b) As shown in Figure 6.9 the use of shafts as the only access shows a decreasing trend since 1970.

(c) As shown in Figure 6.9 the use of declines as the only access shows an increasing trend since 1970.


26

0 5 10 15 20 25 30 35 40

BENCH HEIGHT, hb (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

REL

ATI

VE F

REQ

UEN

CY

SINGLE BENCHES DOUBLE BENCHES

Figure 6.1: Histogram showing the relative frequency of different bench heights, for single and

double benches in open pit mines.

0 50 100 150 200 250 300 350 400

INTERRAMP SLOPE HEIGHT, hr (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

REL

ATI

VE F

REQ

UEN

CY

Figure 6.2: Histogram showing the relative frequency of different interramp slope heights in open

pit mines.


27

0 100 200 300 400 500 600 700 800 900 1000

OVERALL SLOPE HEIGHT, ho (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

REL

ATI

VE F

REQ

UEN

CY

Figure 6.3: Histogram showing the relative frequency of different overall slope heights in open pit mines.

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80 85 90

BENCH FACE INCLINATION, b (degrees)0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

REL

ATI

VE F

REQ

UEN

CY

Figure 6.4: Histogram showing the relative frequency of different bench face inclinations in open

pit mines.


28

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80 85 90

INTERRAMP SLOPE ANGLE, r (degrees)0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

REL

ATI

VE F

REQ

UEN

CY

Figure 6.5: Histogram showing the relative frequency of different interramp slope angles in open

pit mines.

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80 85 90

OVERALL SLOPE ANGLE, o (degrees)0.00

0.05

0.10

0.15

0.20

0.25

0.30

REL

ATI

VE F

REQ

UEN

CY

Figure 6.6: Histogram showing the relative frequency of different overall slope angles in open pit

mines.


29

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80 85 90

SLOPE ANGLE, (degrees)0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

REL

ATI

VE F

REQ

UEN

CY

BENCHES INTERRAMP SLOPES OVERALL SLOPES

Figure 6.7: Histogram showing the relative frequency of different slope angles for benches, inter-

ramp and overall slopes in open pit mines.

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80 85 90

SLOPE ANGLE, (degrees)0

50

100

150

200

250

300

350

400

450

500

550

600

650

700

750

800

850

900

SLO

PE H

EIG

HT,

h

(m)

INTERRAMP SLOPES OVERALL SLOPES

Figure 6.8: Variation of the slope angle with the slope height in open pit mines.


30

(d) Before 1970 in 70% of the cases shafts were used as accesses, in 30% declines were

used, and in 0% both, shafts and declines, were used.

(e) In the period from 1970 to 1990, in 46% of the cases shafts were used as accesses, in 42% declines were used, and in 13% both, shafts and declines, were used.

(f) In the period from 1990 to 2002, in 36% of the cases shafts were used as accesses, in 50% declines were used, and in 14% both, shafts and declines, were used.

6.3. BLOCK HEIGHT AND FOOTPRINT

The analysis of the data on block heights and footprints indicates that:

(a) As shown in Figure 6.10, since 1970 the block height in block/panel caving mines shows an increasing trend. Before 1970, the typical block height was 100 m; for the period 1970-1990 was 160 m, and for the period 1990-2002 it is 240 m.

(b) As shown in Figure 6.11, in block/panel caving mines the footprint area varies widely, but in 80% of the cases, it is smaller than 250000 m2, and its average is 165000 m2.

(c) As shown in Figure 6.12, the footprint geometry is such that the ratio between its length (L) and its width (B) rarely exceeds 3, and in almost 60% of the cases is smaller than 2.

(d) It seems that most block/panel caving mines have ignored a possible relationship be-tween block height (H) and footprint geometry (defined by its width B). As a prelimi-nary conclusion, and as shown in Figure 6.13, the data collected suggested that: o If H/B 1 then the cave will easily connect to surface (or upper level previ-

ously mined out). o If 2 H/B > 1 then the cave probably will connect to surface (or upper level

previously mined out).

TIME PERIOD

0.0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

REL

ATI

VE F

REQ

UEN

CY

ACCESS TYPE

SHAFTS DECLINES BOTH TYPES OF ACCESS

Before 1970 From 1970 to 1990 After 1990

DECLINE

S TREN

D

SHAFTS TREND

BOTHS TREND

Figure 6.9: Evolution through time of the trend for the type of access to underground mines.


31

0 50 100 150 200 250 300 350 400 450 500

BLOCK HEIGHT (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

REL

ATI

VE F

REQ

UEN

CY

TIME PERIOD


Figure 6.10: Evolution through time of the trend for the block height in block/panel caving mines.

0 100000 200000 300000 400000 500000 600000 700000 800000 900000 1000000

FOOTPRINT AREA (m2)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

REL

ATI

VE F

REQ

UEN

CY

AVERAGE FOOTPRINT AREA = 165000 m2

Figure 6.11: Relative frequency of the different footprint area ranges in mines by block/panel caving.


32

o If H/B > 2 then the cave would have problems to connect to surface (or up-per level previously mined out).

Due to the importance of this issue, it will be studied with more accuracy during the development of the geotechnical guidelines that are considered as the second main activity of Task 4.

6.4. CAVING INITIATION

The analysis of the data on caving initiation indicates that:

(a) As shown in Figure 6.14, the shape of the initial area for caving is predominantly square or rectangular, but in a few cases other shapes have been used (like triangular shapes).

(b) As shown in Figure 6.14, the available data indicates that the area for caving initiation has an average value of 10000 m2, and typically varies form 5000 to 15000 m2.

(c) As shown in Figure 6.15, the hydraulic radius of the initial caving area varies from 15 to 45 m, with an average value in the range from 20 to 30 m.

(d) As shown in Figure 6.16, to facilitate cave initiation in 53% of the cases slots have been used, in 7% of the cases artificial chimneys have been used (chimneying inten-tionally used to initiate caving, and not a product of poor cave management), and in 40% of the cases no measures to facilitate cave initiation have been used.

0 100 200 300 400 500 600 700 800 900 1000 1100 1200

FOOTPRINT LENGTH (m)

0

50

100

150

200

250

300

350

400

450

500

550

600

650

700

750

800

FOO

TPR

INT

WID

TH

(m)

L / W = 1.0

5.0

1.5

2.0

2.5

3.0

3.5

4.04.5

0%

CUMULATIVE FREQUENCY

31%

59%

81%

94%

100%

Figure 6.12: Trend for the ratio between the footprint length (L) and its width (B) block/panel caving

mines.


33

0 50 100 150 200 250 300 350 400 450 500 550 600 650 700 750 800

FOOTPRINT WIDTH (m)

0

50

100

150

200

250

300

350

400

450

500

BLO

CK

HE

IGH

T

(m)

DIFFICULTCONNECTION

TO SURFACE ?

EASYCONNECTION TO SURFACE

H = 2B H = B

CONNECTION TO SURFACE

Figure 6.13: Trend between the block height (H) and the footprint width (B) for block/panel caving

mines.

0 5000 10000 15000 20000 25000 30000 35000 40000

INITIAL CAVING AREA (m2)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

REL

ATI

VE F

REQ

UEN

CY

AREA SHAPE

SQUARE RECTANGULAR OTHER

Figure 6.14: Relative frequency of different initial caving areas and their shapes in block/panel

caving mines.


34

0 5 10 15 20 25 30 35 40 45 50

HYDRAULIC RADIUS OF INITIAL CAVING AREA (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

REL

ATI

VE F

REQ

UEN

CY

Figure 6.15: Relative frequency of different hydraulic radius for the initial caving area in

block/panel caving mines.

MEASURES TO FACILITATE CAVING INITIATION

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

REL

ATI

VE F

REQ

UEN

CY

NONESLOT ARTIFICIAL CHIMNEY

Figure 6.16: Relative frequency of different measures to facilitate caving initiation in block/panel

caving mines.


35

6.5. UNDERCUT LEVEL

The analysis of the data on the undercut level indicates that:

(a) As shown in Figure 6.17, the distance between undercut drifts varies from 10 to 35 m, with an average from 20 to 25 m.

(b) As shown in Figure 6.18, the width of undercut drifts shows an increasing trend through time. Before 1970 it has an average from 2 to 3 m, in the period 1970-1990 its average was 3 m, and in the period 1990-2002 its average is 4 m.

(c) As shown in Figure 6.19, the height of undercut drifts shows an increasing trend through time. Before 1970 it has an average from 2.0 to 2.5 m, in the period 1970-1990 its average was 3.0 to 3.5 m, and in the period 1990-2002 its average is from 3.5 to 4.0 m.

(d) As shown in Figure 6.20, the undercut height shows no time-dependent trends. It var-ies from 3 to 20 m, with an average from 8 to 12 m.

(e) As shown in Figure 6.21, the undercut rate varies from 500 to 5000 m2/month, with an average from 2000 to 2500 m2/month.

6.6. EXTRACTION LEVEL

The analysis of the data on the extraction level indicates that:

(a) As shown in Figure 6.22, the nominal crown-pillar thickness (from floor extraction level to floor undercut level) shows an increasing trend through time. Before 1970 its average was from 7.5 to 10.0 m, in the period 1970-1990 it was 12.5, and in the period 1990-2002 it is from 15.0 to 17.5 m.

(b) As shown in Figure 6.23, the spacing between extraction level drifts shows an in-creasing trend through time. Before 1970 its average was from 12 to 16 m, in the pe-riod 1970-1990 it was from 20 to 24 m, and in the period 1990-2002 it is from 26 to 28 m.

(c) As shown in Figure 6.24, the width of extraction level drifts shows an increasing trend through time. Before 1970 it has an average of 2.5 m. In the period 1970-1990 its av-erage was from 3.0 to 3.5 m, and in the period 1990-2002 its average is from 4.0 to 4.5 m.

(d) As shown in Figure 6.25, the height of extraction level drifts shows an increasing trend through time. Before 1970 it has an average from 2.0 to 2.5 m. In the period 1970-1990 its average was 3.0 to 3.5 m, and in the period 1990-2002 its average is from 3.5 to 4.5 m.

(e) As shown in Figure 6.26, the draw point spacing shows an increasing trend through time. Before 1970 it has an average of 8 m. In the period 1970-1990 its average was 12 m, and in the period 1990-2002 its average is 15 m.

(f) As shown in Figure 6.27, the influence area of draw points shows an increasing trend through time. Before 1970 it has an average of 50 m2. In the period 1970-1990 its average was 125 m2, and in the period 1990-2002 its average is from 200 to 225 m2.

(g) As shown in Figure 6.28, the most used geometry for the extraction level is the herringbone layout (54% of the cases), followed by El Teniente layout (layout 40% of the cases).

(h) As shown in Figure 6.29, the average draw rate is from 0.20 to 0.25 m/day.

(i) The current practice is to use draw rates that increase with the percentage of block extraction, as shown in Figure 6.30.


36

0 5 10 15 20 25 30 35

NOMINAL DISTANCE BETWEEN DRIFTS UCL (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

RE

LATI

VE

FR

EQ

UEN

CY

Figure 6.17: Relative frequency of different nominal distances between undercut level drifts in

caving mines.

0 1 2 3 4 5 6 7 8

NOMINAL WIDTH DRIFTS UCL (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

0.75

0.80

0.85

0.90

RE

LATI

VE

FR

EQ

UEN

CY

TIME PERIOD


Figure 6.18: Time trend of the relative frequency for the nominal width of undercut level drifts in

caving mines.


37

0 1 2 3 4 5 6

NOMINAL HEIGHT DRIFTS UCL (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

0.75

0.80

RE

LATI

VE

FR

EQ

UEN

CY

TIME PERIOD


Figure 6.19: Time trend of the relative frequency for the nominal height of undercut level drifts in

caving mines.

0 4 8 12 16 20

UNDERCUT HEIGHT (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

REL

ATI

VE F

REQ

UEN

CY

TIME PERIOD


Figure 6.20: Relative frequency of different undercut heights in caving mines.


38

0 500 1000 1500 2000 2500 3000 3500 4000 4500 5000 5500

AVERAGE UNDERCUT RATE (m2/month)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

REL

ATI

VE F

RE

QU

EN

CY

Figure 6.21: Relative frequency of different undercut rates in caving mines.

0.0 2.5 5.0 7.5 10.0 12.5 15.0 17.5 20.0 22.5 25.0 27.5 30.0 32.5 35.0

NOMINAL CROWN-PILLAR THICKNESS (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

0.75

0.80

0.85

0.90

RE

LATI

VE

FR

EQ

UEN

CY

TIME PERIOD


Figure 6.22: Evolution through time of the trend for nominal crown-pillar thickness in mines by

caving methods.


39

0 4 8 12 16 20 24 28 32 36 40

PRODUCTION DRIFTS SPACING (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

RE

LATI

VE

FR

EQ

UEN

CY

TIME PERIOD


Figure 6.23: Evolution through time of the trend for production drifts spacing in mines by caving.

0 1 2 3 4 5 6 7 8

NOMINAL WIDTH DRIFTS EXTRACTION LEVEL (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

RE

LATI

VE

FR

EQ

UEN

CY

TIME PERIOD


Figure 6.24: Evolution through time of the trend for the width of extraction level drifts in mines by

caving methods.


40

0 1 2 3 4 5 6

NOMINAL HEIGHT DRIFTS EXTRACTION LEVEL (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

RE

LATI

VE

FR

EQ

UEN

CY

TIME PERIOD


Figure 6.25: Evolution through time of the trend for the height of extraction level drifts in mines by

caving methods.

0 2 4 6 8 10 12 14 16 18 20 22 24

DRAW POINT SPACING (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

RE

LATI

VE

FR

EQ

UEN

CY

TIME PERIOD Before 1970 From 1970 to 1990 After 1990

Figure 6.26: Evolution through time of the trend for draw point spacing in mines by block and pa-

nel caving methods.


41

0 25 50 75 100 125 150 175 200 225 250 275 300 325 350

INFLUENCE AREA OF DRAW POINTS (m2)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

RE

LATI

VE

FR

EQ

UEN

CY

TIME PERIOD Before 1970 From 1970 to 1990 After 1990

Figure 6.27: Evolution through time of the trend for the influence area of draw points in mines by

block and panel caving methods.

EXTRACTION LEVEL LAYOUT

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

RE

LATI

VE

FR

EQ

UEN

CY

HERRINGBONE TENIENTE OTHER

Figure 6.28: Relative frequency of different extraction level layouts in mines by block and panel

caving methods.


42

0.00 0.05 0.10 0.15 0.20 0.25 0.30 0.35 0.40 0.45 0.50 0.55 0.60 0.65 0.70

AVERAGE DRAW RATE (m/day)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

REL

ATI

VE F

RE

QU

EN

CY

Figure 6.29: Relative frequency of different average draw rates in mines by caving methods.

0 10 20 30 40 50 60 70

BLOCK EXTRACTION (%)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

DR

AW

RA

TE

(m/d

ay)

Pilar Sub 6 - Esmeralda Sector Hw / Central, Initial Caving Pilar Sub 6 - Esmeralda Sector Fw, Initial Caving Esmeralda, Initial Caving Diablo-Regimiento Project, Initial Caving Palabora, Initial Caving Average for Initial Caving El Teniente trend for Steady-State Caving

Figure 6.30: Examples of the variation of the draw rate as a function of the percentage of block

extraction, in mines by block/panel caving.


43

6.7. SUPPORT

The analysis of the data on support indicates that:

(a) In most underground mines by caving the support at the undercut level includes only bolts; nevertheless, in some mines this support also included mesh and shotcrete.

(b) In most underground mines by caving the support at the extraction level includes bolts (typically from 1.8 to 2.4 m long, at spacings from 1.0 to 1.3 m), mesh and shotcrete (typically 2), and in many cases also cables (typically at intersections, with lengths from 5 to 8 m). Also some mines used straps and osro-straps, as shown in Photo-graph 6.1.

(c) As shown in Figure 6.31, the bolt length varies from 1.25 to 3.75 m, with an average from 2.00 to 2.25 m, for the Undercut Level, and from 2.00 to 2.50 m for the Extraction Level.

(d) As shown in Figure 6.32, the bolt spacing varies from 0.6 to 1.40 m, being typically 1.0 m for both: Undercut and Extraction Levels (50% of cases). The average bolt spacing is from 1.0 to 1.1 m, also for both levels.

(e) The variation of bolt lengths with the width of the drifts is shown in Figure 6.33, which indicates that: o There is no clear difference between the Undercut and Extraction Levels. o In most cases the bolt length is such that: 1.5 B / L 3.0 o For preliminary estimations of bolt length, the following relationships are sug-

gested (the drift width, B, expressed in m): For poor quality rock masses (20 RMR 40): L (m) = 0.60 B + 0.60 For fair quality rock masses (40 RMR 60): L (m) = 0.45 B + 0.45 For good quality rock masses (60 RMR 80): L (m) = 0.30 B + 0.30

(f) The variation of bolt spacing (s) with the bolt length (L) is shown in Figure 6.34, which indicates that:: o There is no clear difference between the Undercut and Extraction Levels. o In most cases the bolt length is such that: 1.5 L / s 2.5

Photograph 6.1: Extraction level support by bolts, mesh and osro-straps at a South African under-ground mine.


44

0.0 0.5 1.0 1.5 2.0 2.5 3.0 3.5 4.0

BOLT LENGTH (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

REL

ATI

VE F

REQ

UEN

CY

UNDERCUT LEVEL EXTRACTION LEVEL Fit 1: Normal

Figure 6.31: Relative frequency of different bolt lengths in mines by caving methods.

0.0 0.2 0.4 0.6 0.8 1.0 1.2 1.4 1.6 1.8 2.0

BOLT SPACING (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

RE

LATI

VE

FR

EQ

UEN

CY

UNDERCUT LEVEL EXTRACTION LEVEL Fit 1: Normal

Figure 6.32: Relative frequency of different bolt spacings in mines by caving methods.


45

3.0 3.5 4.0 4.5 5.0 5.5 6.0 6.5

DRIFT WIDTH, B (m)

1.0

1.5

2.0

2.5

3.0

3.5

4.0

BO

LT L

ENG

TH,

L (

m)

2.0

2.5

3.0

3.5

4.0

4.55.0

B / L = 1.0 1.5

GOOD ROCK

MASS QUAL

ITY, L = 0,3

0 B + 0,30

POOR

ROCK

MASS

QUALITY

, L =

0.60 B +

0.60

FAIR ROC

K MASS Q

UALITY,

L = 0.45

B + 0.45

UNDERCUT LEVEL DATA

EXTRACTION LEVEL DATA

BEST FIT FOR B < 5 m

Figure 6.33: Variation of the bolt length with the nominal width of the drift in mines by caving

methods.

1.0 1.2 1.4 1.6 1.8 2.0 2.2 2.4 2.6 2.8 3.0 3.2 3.4 3.6 3.8 4.0

BOLT LENGTH, L (m)

0.5

0.6

0.7

0.8

0.9

1.0

1.1

1.2

1.3

1.4

1.5

BO

LT S

PA

CIN

G,

s (

m)

UNDERCUT LEVEL DATA

EXTRACTION LEVEL DATA

3.0

3.5

4.0

4.5

5.0

L / s = 1.0 1.5 2.0 2.5

GOOD RO

CK MASS

QUALITY

POOR ROCK

MASS QUALIT

Y

FAIR ROCK

MASS QUAL

ITY

s = 0.150

L + 0.9

s = 0.100 L +

0.7

s = 0.125 L

+ 0.8

Figure 6.34: Variation of the bolt spacing with the bolt length in mines by caving methods.


46

o For preliminary estimations of bolt spacing, the following relationships are sug-gested (the bolt length, L, expressed in m): For poor quality rock masses (20 RMR 40): s (m) = 0.100 L + 0.7 For fair quality rock masses (40 RMR 60): s (m) = 0.125 B + 0.8 For good quality rock masses (60 RMR 80): s (m) = 0.150 B + 0.9

(g) Underground mines by caving methods and under rockburst risk, have also used mesh and lacing as a complementary support for extraction level drifts.

(h) The support of the draw points changes from one mine to another, but in most cases it includes steel arches, cablebolts and concrete and/or shotcrete. The number of steel arches had varied from 2 to 7, but currently most mines used 2 to 3 steel arches. Pho-tographs 6.2 and 6.3 show some examples of draw point support.

Photograph 6.2: Draw point support using steel sets and concrete at a North American mine by caving

Photograph 6.3: Draw point support using steel sets and concrete at a South African mine by caving


47

6.8. MATERIAL HANDLING SYSTEM

The analysis of the data on material handling systems indicates that:

(a) As shown in Figure 6.35, in 57% of the cases underground mines by caving use pro-duction shafts; in 27% of the cases they use conveyor belts; in 12% of the cases they use trains; and in 4% of the cases they use trucks.

MATERIAL HANDLING SYSTEM

0.0

0.1

0.2

0.3

0.4

0.5

0.6

REL

ATI

VE F

REQ

UEN

CY

SHAFTS CONVEYOR BELTS TRUCKS TRAINS

Figure 6.35: Relative frequency of different material handling systems used in underground mines

by caving methods.


48

7. MINE OPERATION DATA

The mine operation data collected include: Operational parameters for the undercut and extraction levels. Production blasting Fragmentation Oversize limits Draw rates Equipment Repair frequencies All the data obtained for each mine visited are included in Appendix B.

The analysis of the mine operation data indicates that:

(a) As shown in Figure 7.1, the powder factor for undercut blasting varies widely, from 200 to 1000 grm/ton; with an average from 400 to 500 grm/ton, and a typical or most used value from 300 to 600 grm/ton.

(b) As shown in Figure 7.2, the LHD capacity varies from 7 to 19 tons, with an average of 11 tons.

(c) As shown in Figure 7.3, the LHD tramming distance varies widely, from 25 to 300 m, with an average from 125 to 150 m.

(d) As shown in Figure 7.4, the oversize limit in most cases (almost 50%) varies form 1.8 to 2.0 m3; nevertheless, its range is wide, from 0.4 to 2.4 m3. The average oversize limit is 1.6 m3.

0 100 200 300 400 500 600 700 800 900 1000 1100 1200

POWDER FACTOR (grm/ton)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

REL

ATI

VE F

REQ

UEN

CY

Figure 7.1: Relative frequency of different values of the powder factor used for undercut blasting

in mines by caving methods.


49

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

LHD CAPACITY (tons)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

REL

ATI

VE F

RE

QU

EN

CY

Figure 7.2: Relative frequency of different LHD capacities used in mines by caving methods.

0 25 50 75 100 125 150 175 200 225 250 275 300

LHD TRAMMING DISTANCE (m)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

REL

ATI

VE F

RE

QU

EN

CY

Figure 7.3: Relative frequency of different LHD tramming distances used in mines by caving

methods.


50

0.0 0.2 0.4 0.6 0.8 1.0 1.2 1.4 1.6 1.8 2.0 2.2 2.4 2.6 2.8 3.0

OVERSIZE LIMIT (m3)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

RE

LATI

VE

FR

EQ

UEN

CY

Figure 7.4: Relative frequency of different oversize limits in mines by caving methods.


51

8. GEOTECHNICAL INSTRUMENTATION AND MONITORING DATA

The data on geotechnical instrumentation and monitoring include: Parameters to be monitored. Purpose. Instruments. Number. Length. Frequency of readings. Threshold values. Degree of satisfaction. All the data obtained for each mine visited are included in Appendix B.

The analysis of the geotechnical instrumentation and monitoring data indicates that:

(a) As shown in Figure 8.1, in open pit mines the frequency of use and degree of satisfaction (DS) with monitoring is:

Degree of Satisfaction Frequency of Use Monitoring System

Range Average

Field inspections (100%) Fair to Very High High Most used

Global displacements (100%) Fair to Very High Fair to High

Local displacements (78%) Very Low to Very High Fair Second most used

Groundwater monitoring (67%) Fair to Very High Fair

Aerial photography (44%) High High Third most used

TDR (33%) Very Low to Fair Fair

(b) As shown in Figure 8.2, in underground mines by caving methods the frequency of use and degree of satisfaction (DS) with monitoring is:

Degree of Satisfaction Frequency of Use Monitoring System

Range Average

Field inspections (100%) Low to High High Most used

Local displacements (82%) Very Low to Very High Fair

Seismic System (64%) Fair to Very High High Second most used

TDR (64%) Low to Very High High

Convergence (36%) High to Very High High Third most used

Observation Boreholes (36%) Low to Very High Fair


52

DEGREE OF SATISFACTION

GEO

TEC

HN

ICA

L M

ON

ITO

RIN

G I

N O

PEN

PIT

MIN

ING

VERY LOW VERY HIGHHIGHFAIRLOW

FIELD INSPECTIONS

AERIAL PHOTOGRAPHS

RECONCILIATION

TIME DOMAIN REFLECTOMETER

GROUNDWATER

GLOBAL DISPLACEMENTS

LOCAL DISPLACEMENTS

FREQUENCYOF USE

78%

100%

67%

33%

11%

44%

100%

Figure 8.1: Relative frequency and degree of satisfaction for different geotechnical instrumenta-tion and monitoring systems used in open pit mines.

DEGREE OF SATISFACTION

GEO

TEC

HN

ICA

L M

ON

ITO

RIN

G I

N U

ND

ERG

RO

UN

D M

ININ

G VERY LOW VERY HIGHHIGHFAIRLOW

CONVERGENCE

FIELD INSPECTIONS

AERIAL PHOTOGRAPHS(SUBSIDENCE)

SEISMIC SYSTEMS

OBSERVATION BOREHOLES(CAVE BACK)

OVERBREAK

TIME DOMAIN REFLECTOMETER(CAVE BACK)

WATER FLOW

STRESSES

GLOBAL DISPLACEMENTS

LOCAL DISPLACEMENTS

FREQUENCYOF USE

36%

82%

9%

27%

18%

18%

64%

36%

64%

27%

100%

Figure 8.2: Relative frequency and degree of satisfaction for different geotechnical instrumenta-tion and monitoring systems used in underground mines by caving methods.


53

9. GEOTECHNICAL HAZARDS DATA

The geotechnical hazards considered in the bench marking includes:

Open pit hazards (rock falls, wedge/planar failures and slides). Underground mines hazards (rib pillar failures, pillar instabilities, stope instabilities, early di-

lution, water inflows and mudrushes, collapses, hangups, rockbursts and subsidence). All the data obtained for each mine visited are included in Appendix B.

9.1. COLLAPSES

A collapse is a type of hazards that frequently affects the extraction level of underground mines by caving methods, causing important damage not only at the undercut level but also at the extraction level, as illustrated by the example shown in Photographs 9.1 and 9.2.

The analysis of the data on collapses indicates that:

(a) As shown in Figure 9.1 the area affected by a single collapse varies from 140 to 17500 m2, with an average of 3700 m2.

(b) As shown in Figure 9.2 the main causes of collapses are: o Draw rate / Draw management o Structures o Mine planning / Mining sequence

(c) As shown in Figure 9.3 the most frequent remedial measures for collapses are: o Draw rate / Draw management o Support o Improving geological-geotechnical data

1,5 m

Photograph 9.1: Collapse at an undercut level drift of Teniente 4 Sur (1989).


54

CONCRETEDAMAGE

CONCRETEDAMAGE

1,5 m

Photograph 9.2: Collapse at an extraction level drift of Teniente 4 Sur (1989).

0 2000 4000 6000 8000 10000 12000 14000 16000 18000 20000

AREA AFFECTED BY A SINGLE COLLAPSE (m2)

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

REL

ATI

VE F

RE

QU

EN

CY

Figure 9.1: Relative frequency of the area affected by a single collapse in underground mines by

caving.


55

MAIN CAUSES OF A COLLAPSE

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

0.75

0.80

0.85

0.90

0.95

1.00

RE

LATI

VE

FR

EQU

ENC

Y

FIRST MORE IMPORTANT CAUSE SECOND MORE IMPORTANT CAUSE THIRD MORE IMPORTANT CAUSE

GEOLOGYSTRUCTURES WATER

MINE LAYOUTDESIGN SUPPORT

MINE PLANNINGMINING

SEQUENCE

DRAW RATEDRAW

MANAGEMENTBLASTING

Figure 9.2: Relative frequency of the different main causes of collapses in mines by caving.

MAIN REMEDIAL MEASURES FOR A COLLAPSE

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0.55

0.60

0.65

0.70

0.75

0.80

0.85

0.90

0.95

1.00

REL

ATI

VE F

REQ

UEN

CY

FIRST MOST COMMON REMEDIAL MEASURE SECOND MOST COMMON REMEDIAL MEASURE THIRD MOST COMMON REMEDIAL MEASURE

IMPROVEDGEOLOGICAL

GEOTECHNICALDATA

DRAINAGE SUPPORT MINE PLANNINGMINING

SEQUENCE

DRAW RATEDRAW

MANAGEMENT

CONTROLLED BLASTING

Figure 9.3: Relative frequency of the different remedial measures against collapses that have

been used in underground mines by caving.


56

9.2. ROCKBURSTS

A rockburst is a seismic event that causes damage. In most cases the damage has no clear structural control, as shown in Picture 9.3, but in certain cases like the one shown in Picture 9.4, the damage has a clear structural control because the seismic event triggered the fall of blocks. The intensity of this damage can vary widely, but for the purposes of this report it will be considered that a rockburst can produce three levels of damage: heavy, moderated, and light damage. These classes of damage are illustrated by the examples shown in Pictures 9.5 to 9.7.

Photograph 9.4: Typical major rockburst damage, with structural control, due to a seismic event that affected a drift at the ventilation level of Teniente Sub 6 (1990).

Photograph 9.3: Typical major rockburst damage, without structural control, due to a seismic

event that affected a drift at the undercut level of Teniente Sub 6 (1991).


57

Photograph 9.5: Example of MODERATE rockburst damage.

Photograph 9.5: Example of HEAVY rockburst damage.


58

Photograph 9.7: Example of LIGHT rockburst damage.

Due to the fact that not all mines by caving suffer rockbursts, and considering that most of the data collected came from El Teniente Sub 6 experience, the numerical conclusions pre-sented below are based on the analysis of these data. The analysis of the data on rock-bursts indicates that:

(a) Rockburst can affect not only the undercut level, but also different levels below the UCL, reaching up to the haulage level.

(b) The major rockbursts that damaged Teniente Sub 6 caused different kinds of damage at different levels, and at different distances form the caving front.

(c) As shown in Figure 9.4 the heavy damage at different levels varies with the distance to the caving front as follows:

Level Distance to Caving Front of Damaged Zone Most Damaged Sector

Undercut 0 to 150 m 0 to 50 m

Extraction < 0 to 150 m 0 to 50 m

Ventilation 50 to 150 m 100 to 150 m

Haulage < 0 to 150 m 100 to 150 m


59

(d) As shown in Figure 9.5 the moderate damage at different levels varies with the dis-tance to the caving front as follows:




Ventilation < 0 to 150 m < 0 to 150 m

Haulage < 0 to 150 m 0 to 50 m

(e) As shown in Figure 9.6 the light damage at different levels varies with the distance to

the caving front as follows:




Ventilation 50 to 150 m 100 to 150 m

Haulage < 0 to 150 m 0 to 50 m

(f) As shown in Figure 9.7 the ma

benchmarking final report

Documents

geotechnical data

available data

general data

data processing

design data

collected data

monitoring data

inclusion of data