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 Karzulovic

December 2002

CONTENTS 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

CONTENTS 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 transition from open pit to underground cave mining. To achieve this objective, the following main activities have been considered: Benchmarking, Geotechnical Guidelines, Guidelines, worked Example 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, BENCHMARKING, 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 collection, an Excel© spreadsheet was designed, and e-mailed to the targeted mines also willing to provide information.

2.

MINE VISITING: VISITING : 17 mines were selected to be visited and relevant information was obtained. 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 underground) that could provide relevant information.

3.

ADDITIONAL DATA COLLECTION : a comprehensive survey of the available technical literature was done in order to collect collect supplementary data. This allows the inclusion 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 PROCESSING: The collected data was analyzed in order to develop histograms and, where possible, correlations showing the current practices and tr ends of underground mining by caving methods. methods. When enough data was available available the relative frequency 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: REPORT : All of the above mentioned, and the conclusions and recommendations 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

Rock Mass Quality

50 ≤ RMR < 60

Acces

Decline

Block Height < 50000 m Footprint Area

50000 a 100000 m > 100000 m

Caving Initiation

l e v e L t u c r e d n U

2

This typical block height could vary ± 20%.

30000 m

2

75000 m

2

170000 m

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.

2

2

Area

10000 m

Shape

Square

Measures to Facilitate

Slot

Is highly recommended to facilitate cave initiation.

Hydraulic Radius

20 to 30 m

Avoid being close to the limit in Laubscher’s chart.

Spacing

15 m

Height Width

4m 4m

Could be increased but not decreased.

Undercut Height

8m

Could vary, but be careful if using small undercutting heights.

Undercut Rate

2100 m /month

s t f i r D

o l i t e c v a e r L t x n E

2

If RMR > 60 rock mass cavability must be evaluated carefully. Currently 70% of mines prefer declines, and 20% declines and shafts as mine access.

210 m 2

Comments

Smaller areas are not recommended, specially in massive rock masses. Internal corners must be avoided (e.g. a “L” shaped area).

This is the current practice.

2

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 Width

4m 4m

s t f i r D

Spacing

Draw Points

Influence Area Draw Rates

LHD Equipment


15 m 225 m

Could vary from 13 to 18 m. 2

2

Could vary from 169 to 324 m . 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.

0.20 m/day

Capacity

11 ton

It could vary ± 20%.

Traming Distance

140 m

Smaller tramming distances are preferable.

Powder Factor

400 grm/ton

Oversize Limit

3

Subsidence

RMR < 70 RMR > 70

1.8 to 2.0 m α >

45° α > 60°

For undercutting blasting. It could vary ± 20%. It could vary ± 20%. α 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 recommended 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 Laubscher’s 1990 system.

ICS-II, Task 4

1.

GEOTECHNICAL GUIDELINES FOR A TRANSITION FROM OPEN PIT TO UNDERGROUND MINING

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 transition 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 develop, 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 t he available technical literature.

2.

GEOTECHNICAL GUIDELINES , to develop practical methodologies to deal with the key issues 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 Santiago, only the first activity, BENCHMARKING, has been developed and it is presented in this report.

1

ICS-II, Task 4

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 collection, 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 Appendix B.

7.

MINE VISITING: 17 mines were selected to be visited and relevant information was obtained. 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 underground) 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 literature 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 tr ends 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 processing are included in Appendix C.

10.

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

2

ICS-II, Task 4


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

Ridgeway

Underground mining by sublevel caving. Visited October 2002

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

Salvador

Underground mining by panel caving. Visited July 2002

Australia

Canada

Chile

Indonesia

Grasberg Underground (DOZ) Grasberg Open Pit

South Africa

Sweden

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

Palabora

Mine developing a transition from open pit to underground mining. Open pit mine and underground mining by panel caving. Visited May 2002

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

Henderson

Underground mining by panel caving. Visited June 2002

USA

3

ICS-II, Task 4

3.


DATA PROCESSING The collected data was analyzed in order to develop histograms and, if possible, correlations showing 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 assessed. 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 developed, 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)

4

ICS-II, Task 4


Table 3.1 MINES THAT DEVELOPED, ARE DEVELOPING OR WILL D EVELOP A TRANSITION FROM OPEN PIT TO UNDERGROUND MINING Country

Mine Argyle Diamond Big Bell

Australia

H PIT (m)

Transition Type Open pit UG mining

→ UG

→ Open

pit

mining →

Mount Keith Open pit

Developments

150 a 300

Mining 2006 (?)

1994

UG mining

Mount Isa

Date

1997

146

1967

344

2015 (?)

UG mining

→

Northparkes

100

1993

1997

76

1963

1964

250

1969

1973

150 (?)

1941

1948

Perseverance Craigmont Canada

Kidd Creek

Open pit

UG mining

→

Stobie Williams Chile

Chuquicamata

Open pit

UG mining

→

Mansa Mina

1100

2016 (?)

400 (?)

2014 (?)

Finland

Pyhasalmi

Open pit

→

UG mining

135

1967

Indonesia

Grasberg

Open pit

→

UG mining

1000

2016 (?)

Open pit

→

UG mining

Russia

South Africa

Sweden

Kirovsky

1959

Mir

455

Finsch

423

Koffiefontein

240

Palabora

1994 1979

1981

803

1996

2000

Premier

189

1945

1946 (?)

Thabazimbi

70 a 240

1988 (?)

Venetia

360 (?)

2011 (?)

230

1958

747 a 899

2012 (?)

Open pit

UG mining

→

Kiruna

Open pit

→

Bingham Canyon

Open pit

→ UG

UG mining mining

Climax USA

1973 UG mining

→ Open

pit

Miami San Manuel Questa

Zambia

1990

Nchanga

UG mining

→ Open

Open pit UG mining

→

→ Open

pit

→

UG mining

UG mining pit

→

150 (?)

UG mining

1979

1983

1937

1939

1955

1957

Gaths Zimbabwe

Miriam

Open pit

→ UG

mining

60

Shabanie

150 (?)

1950 (?)

Shangani

150

1980 (?)

5

ICS-II, Task 4


1200 MINES THAT DEVELOPED A TRANSITION FROM OP TO UG

1100

CHUQUICAMATA

MINES THAT ARE DEVELOPING A TRANSITION FROM OP TO UG MINES THAT WILL DEVELOP A TRANSITION FROM OP TO UG

1000

GRASBERG TREND FROM CASE HISTORIES TREND FROM PROJECT DATA

) 900 m (

H T P E D T I P M U M I X A M

BINGHAM CANYON

800

PALABORA

700 600 500 400 300 200 100 0 1945

1950

1955

1960

1965

1970

1975

1980

1985

1990

1995

2000

2005

2010

2015

2020

EAR

Figure 3.1:

Evolution through time of the trend for t he depth of open pit mines that have developed, are developing, or will develop a transition to underground mining.

6

ICS-II, Task 4

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 geotechnical group and the ore production for open pit and underground mining. These trends indicate that: 



(d)

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 geotechnical group when the ore production exceeds 25 kTPD in underground mines, and 75 kTPD, in open pit mines.

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).

7

ICS-II, Task 4


30

) e l p o e p (

G N I N I M ) D D N P U T O R 5 k G 2 R E a n h D t N U r e E m o V I ( S S A M

25

20

P U O R G 15 L A C I N H C 10 E T O E G

N G N I M I ) P I T T P D N 5 k E 7 O P n E t h a V I r e S S m o M A (

UG (< 25 k TPD )

5

OP (< 75 kT PD)

MINING METHOD OPEN PIT UNDERGROUNG

0 0

25

50

75

100

125

150

175

200

225

250

ORE PRODUCTION (kTPD)

Figure 4.1:

Variation of the size of the typical geotechnical group with ore production in open pit and underground mining.

8

ICS-II, Task 4

5.


GEOTECHNICAL DATA The geotechnical data collected included information on Structures, Rock Mass, Stress Environment, 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 (φ 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 predominate (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 cohesion 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 mechanical 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 (γ), 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).

9

ICS-II, Task 4


1

85 - 90

3

2

4

3

1

75 - 79

1

1

2

1

2

1

1

1

4

70 - 74

2

65 - 69

1

1

1

6

1

60 - 64

1

50 - 54

1

45 - 49

2

2 1 1

1

1

3

4

3 4

3

1

1

1

2

1

2

11

3

1

1 1

1

1

1

2

1

3

1

1

1

1

2

1

4

2

1

3

3

1

3

1

1

3

2

2

2 1

1 1

1 2 2

1

55 - 59

D

1

80 - 84

1

2

1 1

1

2

1

1

I P

1

40 - 44

1

35 - 39

1

3

1 1

30 - 34

25 - 29

1

20 - 24

1

15 - 19

10 - 14

1

5 - 9

1

0 - 4

0 -1 9

2 0- 3 9

4 0- 5 9

6 0- 7 9

8 0- 9 9

1 00 - 1 19

1 20 - 1 39

1 40 - 1 59

1 60 - 1 79

D IP

Figure 5.1:

1 80 - 1 99

2 00 - 2 19

2 20 - 23 9

24 0 - 25 9

26 0 - 27 9

2 80 - 29 9

3 00 - 3 19

3 20 - 3 39

3 40 - 3 59

D IR EC TI ON

Trend of the orientation (defined by dip and dip direction) of structural sets in underground mines that use caving methods.

RELATIVE FREQUENCY 0.00

0.02

0.04

0.06

0.08

0.10

0.12

0.14

0.16

0.18

0.20

0.22

0

10

20

30

) s s e r 40 g e d (

P I D

50

60

70

80

90

Figure 5.2:

Histogram showing the relative frequency of different dip angles for structural sets in underground mines that use caving methods.

10

ICS-II, Task 4


200

) a P k ( 175

OPEN PIT MINING UNDERGROUND MINING

S E R U 150 T C U R T 125 S L A C I 100 G O L O 75 E G F O 50 N O I S E 25 H O C 0 15

20

25

30

35

40

45

FRICTION ANGLE OF GEOLOGICAL STRUCTURES (degrees)

Figure 5.3:





Variation of the cohesion of structures with their friction angle, for open pit and underground mines.

Rock mass quality: RQD, RMRBIENIAWSKI , RMRLAUBSCHER , Q, GSI. Rock mass properties: cohesion (c), friction angle (φ ), modulus of deformability (E), Poisson’s ratio (ν), bulk modulus (B), shear modulus (G), wave velocity for P and S waves (VP and VS).

All the data obtained for each mine visited are included in Appendix B. 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 (γ), 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 difference 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 Laubscher’s RMR (53%), followed by Barton’s Q (26%), and Bieniawski’s RMR (15%). The most used method for rock mass classification in open pit mines is Hoek’s GSI (39%), followed by Bieniawski’s RMR (26%), and Laubscher’s RMR (22%). This is shown in Figure 5.6.

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ICS-II, Task 4


0.55

MINING METHOD 0.50

OPEN PIT OPEN STOPING SUBLEVEL CAVING

0.45

BLOCK CAVING PANEL CAVING

0.40

Y C N 0.35 E U Q E 0.30 R F E 0.25 V I T A 0.20 L E R 0.15 0.10

0.05

0.00 0

50

100

150

200

250

300

350

400

450

500

IN TA C T R OC K U NIA XIA L C OMP RE SS I E S TR EN GTH , U C S (MPa)

Figure 5.4:

Relative frequency of the intact rock’s uniaxial compressive strength, UCS, in different mining methods.

0.70 MINING METHOD

0.65

OPEN PIT OPEN STOPING

0.60

SUBLEVEL CAVING BLOCK CAVING

0.55

PANEL CAVING

Y 0.50 C N 0.45 E U Q 0.40 E R F 0.35 E V I 0.30 T A 0.25 L E R 0.20 0.15 0.10 0.05 0.00 0

10

20

30

40

R OC K QU AL IT

Figure 5.5:

50

60

70

80

90

100

D ES IGN ATIO N, R QD (%)

Relative frequency of the Rock Quality Designation Index, RQD, in different mining methods.

12

ICS-II, Task 4


0.55 MINING METHOD 0.50

OPEN PIT UNDERGROUND

0.45 0.40

Y C N 0.35 E U Q E 0.30 R F E 0.25 V I T A 0.20 L E R 0.15 0.10 0.05 0.00 Q (Barton et al.)

RMR (Bieniawski)

RMR (Laubscher)

GSI (Hoek et al.)

ROCK MASS CLASSIFICATION SYSTEM

Figure 5.6:

Methods used in mining for rock mass classification.

(e)

Interpreting all rock mass classification data in terms of Laubscher’s 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

(g)

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

As shown in Figure 5.13, the average trend relating Laubscher´s 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 confinement 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 underground 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 r ock mass properties is, in most cases, poor to fair.

13

ICS-II, Task 4


0.60

MINING METHOD 0.55 OPEN PIT OPEN STOPING

0.50

SUBLEVEL CAVING BLOCK CAVING PANEL CAVING

0.45

Y C 0.40 N E U 0.35 Q E R F 0.30 E V I 0.25 T A L 0.20 E R 0.15 0.10

0.05

0.00 0

10

20

30

40

50

60

70

80

90

100

LAUBSCHER´S ROCK MASS RATING, RMR Figure 5.7: Relative frequency of Laubscher’s Rock Mass Rating, RMR, in open pits and underground mines that use different mining methods.

0.60 0.55

0.50

0.45

Y C 0.40 N E U 0.35 Q E R 0.30 F E V 0.25 I T A L 0.20 E R 0.15

0.10

0.05

0.00 0

10

20

30

40

50

60

70

80

90

100

L AU BS C HE R´S R OC K MA SS R AT IN G, R MR

Figure 5.8:

Relative frequency of Laubscher’s Rock Mass Rating, RMR, in open pit mining.

14

ICS-II, Task 4


0.40

0.35

0.30

Y C N E 0.25 U Q E R 0.20 F E V I T A 0.15 L E R 0.10

0.05

0.00 0

10

20

30

40

50

60

70

80

90

100


Figure 5.9:

Relative frequency of Laubscher’s Rock Mass Rating, RMR, in open stoping mining.

0.40

0.35

0.30


0.05

0.00 0

10

20

30

40

50

60

70

80

90

100

LAUBSCHER´S ROCK MASS RATING, RMR

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

15

ICS-II, Task 4


0.35

0.30

Y 0.25 C N E U Q 0.20 E R F E V I 0.15 T A L E R 0.10

0.05

0.00 0

10

20

30

40

50

60

70

80

90

100


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

0.40

0.35

0.30


0.05

0.00 0

10

20

30

40

50

60

70

80

90

100


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

16

ICS-II, Task 4


MRM R / RMR =

90

1.2

1.1

1 .0

0.9

MINING METHOD OPEN PIT

80

0.8

OPEN STOPING SUBLEVEL CAVING

70

0.7

BLOCK CAVING PANEL CAVING

60

0.6

R50 M R M40

0.5

30

20

10

0 0

10

20

30

40

50

60

70

80

90

100

LAUBSCHER´S RMR

Figure 5.13: Relationship between Laubscher’s Rock Mass and Mining Rock Mass Ratings, RMR and MRMR.

10000

5000

) a P k (

2000

N1000 O I S E 500 H O C S 200 S A M 100 K C O 50 R

OPEN PIT MINING

20

UNDERGROUND MINING

10 15

20

25

30

35

40

45

50

55

60

65

70

R OC K MA SS F RIC TION A NG LE (degrees)

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

17

ICS-II, Task 4


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 (α 1, α 2, α 3), and trends ( β 1, β 2, β 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 (γz). 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)

(g)

≥

KMEAN

≥

0.3 + (100/ z)

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

1.0 + (1500/ z)

≥

KMAX

≥ ≥

0.2 + (100/ z) 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.

18

ICS-II, Task 4


0.35

0.30


0.05

0.00 0

10

20

30

40

50

60

70

80

90

100

MAJOR PRINCIPAL STRESS, S 1 (MPa)

Figure 5.15: Histogram showing the relative frequency of major principal stresses, S1, with different magnitudes (measurements in underground mines).

0.35

0.30


0.05

0.00 0

5

10

15

20

25

30

35

40

MINOR PRINCIPAL STRESS, S 3 (MPa)

Figure 5.16: Histogram showing the relative frequency of major principal stresses, S3, with different magnitudes (measurements in underground mines).

19

ICS-II, Task 4


0.20

0.15

Y C N E U Q E R 0.10 F E V I T A L E R 0.05

0.00 0

5

10

15

20

25

30

35

40

45

50

55

60

65

70

PRINCIPAL STRESS DIFFERENCE, S 1 - S3 (MPa)

Figure 5.17: Histogram showing the relative frequency of major principal stress differences, S1 - S3, with different magnitudes (measurements in underground mines).

VERTICAL STRESS (MPa) 0

5

10

15

20

25

30

35

40

45

50

55

60

65

70

75

80

85

90

0

250 500

750 1000

) s r e 1250 t e m (

H T P E D

1500

1750

2000 2250

2500 2750

3000

Figure 5.18:

Variation of in situ vertical stresses with depth in underground mines.

20

ICS-II, Task 4


K MEAN 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

0

250

500

750

1000

) s r e 1250 t e m ( 1500

H T P 1750 E D 2000 2250

2500 2750

3000

Figure 5.19: Variation of the average value of the in situ stress ratio, KMEAN, with depth in underground mines. The black curves shown the upper and lower boundaries defined by Hoek & Brown (1980), while the red curve is the average between them.

K MIN 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

0

250

500

750

1000

) s r e 1250 t e m ( 1500

H T P 1750 E D 2000

2250

2500

2750

3000

Figure 5.20: Variation of the minimum value of the in situ stress ratio, KMIN, with depth in underground mines. The black curves shown the upper and lower boundaries defined in this work, while the red curve is the average between them.

21

ICS-II, Task 4


K MAX 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

0

250

500

750

1000

) s r e 1250 t e m ( 1500

H T P 1750 E D 2000

2250

2500

2750

3000

Figure 5.21: Variation of the maximum value of the in situ stress ratio, KMAX, with depth in underground mines. The black curves shown the upper and lower boundaries defined in this work, while the red curve is the average between them.

(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 underground 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 software are: SLIDE (30%), DIPS (20%), and SWEDGE (17%).

(b)

As shown in Figure 5.24, for two-dimensional numerical analyses the most used software 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%).

22

ICS-II, Task 4


0.8 UNDERGROUND MINING OPEN PIT MINING 0.7

0.6


0.1

0.0 D RAINA GE TUNNELS

PUM PI NG WELLS

SUBHORIZONTAL DRAI NS

SUMPS

DEWATERING SYSTEM

Figure 5.22: Relative frequency of different dewatering systems used in open pits and underground mines.

0.30 CONVENTIONAL SLOPE STABILITY SOFTWARE CURRENTLY USED IN OPEN PIT MINES

0.25

Y C 0.20 N E U Q E R 0.15 F E V I T A L 0.10 E R

0.05

0.00 SLIDE

DIPS

SWEDGE

XSTABL

ROCFALL

BACKBREAK GALENA

SLOPE/W

UTEXAS

NFOLD

SOFTWARE PACKAGE

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

23

ICS-II, Task 4


0.55 2D NUMERICAL ANALYSIS SOFTWARE CURRENTLY USED IN OPEN PIT AND UNDERGROUND MINES

0.50

0.45

0.40


0.05

0.00

FLAC

UDEC

EXAMINE / EXAMINE TAB

PHASE2

SOFTWARE PACKAGE

Figure 5.24:

Relative frequency of software used in open pit and underground mines for twodimensional numerical analyses.

0.45 3D NUMERICAL ANALYSIS SOFTWARE CURRENTLY USE D IN OPEN PIT AND UNDERGROUND MINES

0.40

0.35

Y C 0.30 N E U Q 0.25 E R F E 0.20 V I T A L 0.15 E R 0.10

0.05

0.00

FLAC3D

3DEC

MAP3D

EXAMINE3D

BEFE

ELAST-3

SOFTWARE PACKAGE

Figure 5.25: Relative frequency of software used in open pit and underground mines for threedimensional numerical analyses.

24

ICS-II, Task 4

6.


MINE DESIGN DATA On open pit mines the mine design data collected included information on Slopes Geometrical Parameters, 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 Systems. All the data obtained for each mine visited are included in Appendix B.

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 doubled). 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 25º to 60º; but in most of the cases it varies from 40° to 60°, and its average is about 50º.

(f)

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

(g)

As shown in Figure 6.7, the slope angle is maximum at bench scale, flatter for interramp 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 estimate 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 t he 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.

25

ICS-II, Task 4


0.70 SINGLE BENCHES

0.65

DOUBLE BENCHES

0.60 0.55


5

10

15

20

25

30

35

40

BENCH HEIGHT, hb (m)

Figure 6.1:

Histogram showing the relative frequency of different bench heights, for single and double benches in open pit mines.

0.45

0.40

0.35


0.05

0.00 0

50

100

150

200

250

300

350

400

INTERRAMP SLOPE HEIGHT, h r (m)

Figure 6.2:

Histogram showing the relative frequency of different interramp slope heights in open pit mines.

26

ICS-II, Task 4


0.35

0.30

Y 0.25 C N E U Q 0.20 E R F E V 0.15 I T A L E R 0.10

0.05

0.00 0

100

200

300

400

500

600

700

800

900

1000

OVERALL SLOPE HEIGHT, h o (m)

Figure 6.3:

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

0.40

0.35

0.30

Y C N E 0.25 U Q E R F 0.20 E V I T A 0.15 L E R 0.10

0.05

0.00 0

5

10

15

20

25

30

35

40

45

50

BENCH FACE INCLINATION,

Figure 6.4:

55

αb

60

65

70

75

80

85

90

(degrees)

Histogram showing the relative frequency of different bench face inclinations in open pit mines.

27

ICS-II, Task 4


0.40

0.35

0.30


0.05

0.00 0

5

10

15

20

25

30

35

40

45

50

INTERRAMP SLOPE ANGLE,

Figure 6.5:

55

60

α r

65

70

75

80

85

90

(degrees)

Histogram showing the relative frequency of different interramp slope angles in open pit mines.

0.30

0.25

Y C 0.20 N E U Q E R F 0.15 E V I T A L 0.10 E R

0.05

0.00 0

5

10

15

20

25

30

35

40

45

50

OVERALL SLOPE ANGLE,

Figure 6.6:

55

αo

60

65

70

75

80

85

90

(degrees)

Histogram showing the relative frequency of different overall slope angles in open pit mines.

28

ICS-II, Task 4


0.40 BENCHES INTERRAMP SLOPES

0.35

OVERALL SLOPES

0.30


0.05

0.00 0

5

10

15

20

25

30

35

40

45

SLOPE ANGLE,

Figure 6.7:

50

α

55

60

65

70

75

80

85

90

(degrees)

Histogram showing the relative frequency of different slope angles for benches, interramp and overall slopes in open pit mines.

900 850

INTERRAMP SLOPES OVERALL SLOPES

800 750 700

) 650 m 600 (

h 550 , T 500 H G 450 I E H 400 E 350 P O 300 L S 250 200 150 100 50 0 0

5

10

15

20

25

30

35

40

45

SLOPE ANGLE,

Figure 6.8:

50

α

55

60

65

70

75

80

85

90

(degrees)

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

29

ICS-II, Task 4


0.8 ACCESS TYPE SHAFTS

0.7

DECLINES BOTH TYPES OF ACCESS

S H A F T ´ S

0.6

Y C N E 0.5 U Q E R F 0.4 E V I T A 0.3 L E R

T R E N D

D E N T R ´ S E I N C L D E

E N D ´ S T R B O T H

0.2

0.1

0.0 Before 1970

From 1970 to 1990

After 1990

TIME PERIOD

Figure 6.9:

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

(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, 2 2 but in 80% of the cases, it is smaller than 250000 m , and its average is 165000 m .

(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 between block height (H) and footprint geometry (defined by its width B). As a preliminary conclusion, and as shown in Figure 6.13, the data collected suggested that: o

o

If H/B ≤ 1 → then the cave will easily connect to surface (or upper level previously mined out). If 2 ≥ H/B > 1 → then the cave probably will connect to surface (or upper level previously mined out).

30

ICS-II, Task 4


0.60

TIME PERIOD

0.55

Before 1970 From 1970 to 1990

0.50

After 1990

0.45

Y C N E U Q E R F E V I T A L E R

0.40

0.35 0.30

0.25 0.20

0.15

0.10 0.05

0.00 0

50

100

150

200

250

300

350

400

450

500

BLOCK HEIGHT (m)

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

0.30

0.25

AVERAGE FOOTP RINT AREA = 165 000 m2


0.05

0.00 0

100000

200000

300000

400000

500000

600000

700000

800000

900000

1000000

2

FOOTPRINT AREA (m )

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

31

ICS-II, Task 4


L / W = 1.0

800

1.5

750 700

0 %

650

2.0

600

) m 550 (

H T D I W T N I R P T O O F

1 %

C 3 U M U L A T I V 5 9 % E F R E % Q U 8 1 E N C Y 9 4 %

500 450 400 350 300 250

2.5

3.0 3.5 4.0 4.5 5.0

0 % 1 0

200 150 100 50 0 0

100

200

300

400

500

600

700

800

900

1000

1100

1200

FOOTPRINT LENGTH (m)

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

o

If H/B > 2 → then the cave would have problems to connect to surface (or upper 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 2 2 has an average value of 10000 m , and typically varies form 5000 to 15000 m .

(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 intentionally 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.

32

ICS-II, Task 4


H = 2B

H=B

500

450

CONNECTION TO SUR FACE

DIFFICULT CONNECTION TO SURFACE ?

400

EASY CONNECTION TO SURFA CE

) 350 m (

T 300 H G I E 250 H K C 200 O L B 150 100

50

0 0

50

100

150

200

250

300

350

400

FOOTPRINT

450

500

550

600

650

700

750

800

IDTH (m)

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

0.35

AREA SHAPE SQUARE

0.30

RECTANGULAR OTHER


0.25

0.20

0.15

0.10

0.05

0.00 0

5000

10000

15000

20000

25000

30000

35000

40000

INIT IA L C A IN G A RE A (m2)

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

33

ICS-II, Task 4


0.40

0.35

0.30


0.05

0.00 0

5

10

15

20

25

30

35

40

45

50

H YD RAU LIC R AD IU S OF INIT IAL C A IN G A RE A (m)

Figure 6.15: Relative frequency of different hydraulic radius for the initial caving area in block/panel caving mines.

0.55

0.50

0.45

0.40

Y C N 0.35 E U Q E 0.30 R F E 0.25 V I T A 0.20 L E R 0.15

0.10

0.05

0.00

SLOT

ARTIFICIAL CHIMNEY

NONE

ME AS UR ES TO FA C ILIT AT E C A IN G IN IT IA TI ON

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

34

ICS-II, Task 4


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 19701990 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 varies 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 m /month, with an 2 average from 2000 to 2500 m /month.

2

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 increasing trend through time. Before 1970 its average was from 12 to 16 m, in the period 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 average 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 2 through time. Before 1970 it has an average of 50 m . In the period 1970-1990 its 2 2 average was 125 m , and in the period 1990-2002 its average is from 200 to 225 m .

(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.

35

ICS-II, Task 4


0.35

0.30

Y 0.25 C N E U Q 0.20 E R F E V 0.15 I T A L E R 0.10

0.05

0.00 0

5

10

15

20

25

30

35

NOMINAL DISTANCE BETWEEN DRIFTS UCL (m)

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

0.90

TIME PERIOD

0.85

Before 1970

0.80

From 1970 to 1990

0.75

After 1990

0.70

Y 0.65 C 0.60 N E 0.55 U Q 0.50 E R F 0.45 E 0.40 V I T 0.35 A L 0.30 E R 0.25 0.20 0.15 0.10 0.05 0.00 0

1

2

3

4

5

6

7

8

NOMINAL WIDTH DRIFTS UCL (m)

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

36

ICS-II, Task 4


0.80

TIME PERIOD

0.75

Before 1970

0.70

From 1970 to 1990 After 1990

0.65 0.60

Y C 0.55 N E 0.50 U Q 0.45 E R F 0.40 E 0.35 V I T A 0.30 L E 0.25 R 0.20 0.15 0.10 0.05 0.00 0

1

2

3

4

5

6

NOMINAL HEIGHT DRIFTS UCL (m)

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

0.45

TIME PERIOD Before 1970

0.40

From 1970 to 1990 After 1990

0.35


0.05

0.00 0

4

8

12

16

20

UNDERCUT HEIGHT (m)

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

37

ICS-II, Task 4


0.30

0.25

Y C 0.20 N E U Q E R 0.15 F E V I T A L 0.10 E R

0.05

0.00 0

500

1000

1500

2000

2500

3000

3500

4000

4500

5000

5500

A ERA GE U ND ERC UT R ATE (m /month) 2

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

0.90

TIME PERIOD

0.85

Before 1970

0.80

From 1970 to 1990

0.75

After 1990

0.70

Y 0.65 C 0.60 N E 0.55 U Q 0.50 E R 0.45 F E 0.40 V I T 0.35 A L 0.30 E R 0.25 0.20 0.15 0.10 0.05 0.00 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)

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

38

ICS-II, Task 4


0.55

TIME PERIOD 0.50

Before 1970 From 1970 to 1990 After 1990

0.45

0.40


0.05

0.00 0

4

8

12

16

20

24

28

32

36

40

PRODUCTION DRIFTS SPACING (m)

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

0.70

TIME PERIOD

0.65


0.60

After 1990

0.55

Y 0.50 C N 0.45 E U Q 0.40 E R F 0.35 E V 0.30 I T A 0.25 L E R 0.20 0.15 0.10 0.05 0.00 0

1

2

3

4

5

6

7

8

N OMIN AL W ID TH D RIF TS E XT RA CT ION L E E L (m)

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

39

ICS-II, Task 4


0.70

TIME PERIOD

0.65


0.60

After 1990

0.55


1

2

3

4

5

6

N OMIN AL H EIG HT D RIF TS E XTR AC TIO N L E E L (m)

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

0.45

TIME PERIOD Before 1970

0.40

From 1970 to 1990 After 1990

0.35


0.05

0.00 0

2

4

6

8

10

12

14

16

18

20

22

24

DRAW POINT SPACING (m)

Figure 6.26: Evolution through time of the trend for draw point spacing in mines by block and panel caving methods.

40

ICS-II, Task 4


0.55

TIME PERIOD 0.50

Before 1970 From 1970 to 1990 After 1990

0.45

0.40


0.05

0.00 0

25

50

75

100

125

150

175

200

225

250

275

300

325

350

IN FL UEN CE A RE A OF D RA W P OIN TS (m2)

Figure 6.27: Evolution through time of the trend for the influence area of draw points in mines by block and panel caving methods.

0.60

0.55

0.50

0.45

Y C 0.40 N E U 0.35 Q E R 0.30 F E V 0.25 I T A L 0.20 E R 0.15

0.10

0.05

0.00

HERRINGBONE

TENIENTE

OTHER

EXTRACTION LE EL LAYOUT

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

41

ICS-II, Task 4


0.40

0.35

0.30

Y C N E 0.25 U Q E R 0.20 F E V I T 0.15 A L E R 0.10

0.05

0.00 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

A ERAGE DRAW RATE (m/day)

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

0.50

0.45

0.40

) 0.35 y a d / m0.30 (

E T 0.25 A R W0.20 A R D 0.15 Pilar Sub 6 - Esmeralda Sector Hw / Central, Initial Caving Pilar Sub 6 - Esmeralda Sector Fw, Initial Caving

0.10

Esmeralda, Initial Caving Diab lo-Regimiento Project, Initial Caving Pala bora, Initial Caving

0.05

A verage for Initial Caving E l Teniente trend for Steady-State Caving

0.00 0

10

20

30

40

50

60

70

BLOCK EXTRACTION (%)

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.

42

ICS-II, Task 4


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 Photograph 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

o

(f)

≤

B / L

≤

3.0

For preliminary estimations of bolt length, the following relationships are suggested (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

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 underground mine.

43

ICS-II, Task 4


0.45 UNDERCUT LEVEL EXTRACTION LEVEL

0.40

Fit 1: Normal 0.35


0.05

0.00 0.0

0.5

1.0

1.5

2.0

2.5

3.0

3.5

4.0

BOLT LENGTH (m)

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

0.50 UNDERCUT LEVEL EXTRACTION LEVEL

0.45

Fit 1: Normal 0.40

Y 0.35 C N E U 0.30 Q E R 0.25 F E V I 0.20 T A L E 0.15 R 0.10

0.05

0.00 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)

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

44

ICS-II, Task 4


B / L = 1.0

4.0

1.5

3.5

O R P O

) m3.0 (

C K R O

S S M A

T Y , A L I Q U

0 L =

.6 0 B

+

0 0 .6

S S M A O C K R R F A I

L , H T G 2.5 N E L T L O 2.0 B

, L L I T Y Q U A

+ . 4 5 B = 0

0. 4 5

2.0

2.5

0, 3 0 B + 0 , 3 0 L = , Y L I T Q UA A S S C K M O R D G O O

3.0

3.5 4.0

1.5

4.5 UNDERCUT LEVEL DATA

5.0

EXTRACTION LEVEL DATA BEST FIT FOR B < 5 m

1.0 3.0

3.5

4.0

4.5

5.0

5.5

6.0

6.5

D RIF T WID TH , B (m)

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

L / s = 1.0

1.5

1.5

2.0

1.4

2.5

D G O O

L I T Y Q UA A S S M K R O C

1.3 F A I R

) 1.2 m (

s , 1.1 G N I C 1.0 A P S 0.9 T L O B 0.8

3.0

3.5

. 9 + 0 0 L 0. 1 5 s =

+ 5 L 0 .1 2 s =

K R O C

L I T Y Q UA MA S S

C K R R O P O O

T Y UA L I S Q M A S

4.0

0. 8

4.5 0. 7 L + 0 . 1 0 0 s =

5.0

0.7

0.6

UNDERCUT LEVEL DATA EXTRACTION LEVEL DATA

0.5 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

B OL T L EN GTH , L (m)

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

45

ICS-II, Task 4


o

For preliminary estimations of bolt spacing, the following relationships are suggested (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 caving methods methods and under under rockburst risk, have have also used mesh and lacing as a complementary support for extraction level drifts.

(h)

The support of the draw points changes changes from one mine mine to another, but but in most cases 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. Photographs 6.2 and 6.2 and 6.3 show 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

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

46

ICS-II, Task 4


0.6

0.5


0.1

0.0 SHAFTS

CONVEYOR BELTS

TRUCKS

TRAINS

MATERIAL HANDLING SYSTEM Figure 6.35: Relative frequency of different different material handling systems used in underground underground mines by caving methods.

6.8. MATERIAL HANDLING SYSTEM The analysis of the data on material handling systems systems indicates that: (a)

As shown in Figure 6.35, 6.35, in 57% of the cases underground mines by caving use production 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.

47

ICS-II, Task 4

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 are included in Appendix B. B. The analysis of the mine operation data indicates that: (a)

As shown in Figure 7.1, 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, 7.2, the LHD capacity varies from 7 to 19 tons, with an average of 11 tons.

(c)

As shown in Figure 7.3, 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, 7.4, the oversize limit in most cases (almost 50%) varies form 1.8 to 2.0 3 3 3 m ; nevertheless, its range is wide, from 0.4 to 2.4 m . The average oversize limit limit is 1.6 m .

0.30

0.25


0.05

0.00 0

100

200

300

400

500

600

700

800

900

1000

1100

1200

POWDE POWDER R FACTO FACTOR (grm/ton)

Figure 7.1:

Relative frequency frequency of different values of the powder factor factor used for undercut undercut blasting in mines by caving methods.

48

ICS-II, Task 4


0.35

0.30


0.05

0.00 0

1

2

3

4

5

6

7

8

9

10

11

12

13

14

15

16

17

18

19

20

LHD CAPACITY (tons)

Figure 7.2:

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

0.35

0.30


0.05

0.00 0

25

50

75

100

125

150

175

200

225

250

275

300

LHD TRAMMING DISTANCE (m)

Figure 7.3:

Relative frequency of different LHD tramming distances used in mines by caving methods.

49

ICS-II, Task 4


0.50

0.45

0.40

Y 0.35 C N E U 0.30 Q E R 0.25 F E V I T 0.20 A L E R 0.15 0.10

0.05

0.00 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

3

O ER S SIIZE LIMIT (m )

Figure 7.4:

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

50

ICS-II, Task 4

8.


GEOTECHNICAL INSTRUMENTATION AND MONITORING DATA The data on geotechnical instrumentation 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 are included in Appendix B. B. The analysis of the geotechnical instrumentation and monitoring data indicates that: (a)

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

Most used

Second most used

Third most used

(b)

Monitoring System

Degree of Satisfaction Range

Average

Field inspections (100%)

Fair to Very High

High

Global displacements (100%)

Fair to Very High

Fair to High

Local displacements (78%)

Very Low to Very High

Fair

Groundwater monitoring (67%)

Fair to Very High

Fair

Aerial photography (44%)

High

High

TDR (33%)

Very Low to Fair

Fair

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

Most used

Second most used

Monitoring System

Degree of Satisfaction Range

Average

Field inspections (100%)

Low to High

High

Local displacements (82%)

Very Low to Very High

Fair

Seismic System (64%)

Fair to Very High

High

TDR (64%)

Low to Very High

High

Convergence (36%)

High to Very High

High

Observation Boreholes (36%)

Low to Very High

Fair

Third most used

51

ICS-II, Task 4


DEGREE DEGREE OF SATIS SATISFAC FACTI TION ON VERY LOW

G N I N I M T I P N E P O N I

LOW

FAIR

HIGH

VERY HIGH

FREQUENCY OF USE

78%

LOCAL DISPLACEMENTS

100%

GLOBAL DISPLACEMENTS

G N I R O T I TIME N O M L A C I N H C E T O E G

GROUNDWATER

67%

DOMAIN REFLECTOMETER

33%

RECONCILIATION

11%

AERIAL PHOTOGRAPHS

44%

Figure 8.1:

100%

FIELD INSPECTIONS

Relative frequency and and degree degree of satisfaction for different geotechnical geotechnical instrumentainstrumentation and monitoring systems used in open pit mines.

DEGREE DEGREE OF SATI SATISFACTI SFACTION ON G N I N I M D N U O R G R E D N U N I

VERY LOW

FAIR

HIGH

VERY HIGH

FREQUENCY OF USE

CONVERGENCE

36%

LOCAL DISPLACEMENTS

82%

GLOBAL DISPLACEMENTS

9%

STRESSES

27%

WATER FLOW

18%

G N I OVERBREAK R O T I TIME DOMAIN REFLECTOMETER (CAVE BACK) N O OBSERVATION BOREHOLES M (CAVE BACK) L A C SEISMIC SYSTEMS I N H AERIAL PHOTOGRAPHS C (SUBSIDENCE) E T O FIELD INSPECTIONS E G

Figure 8.2:

LOW

18% 64% 36% 64% 27% 100%

Relative frequency and and degree degree of satisfaction for different geotechnical geotechnical instrumentainstrumentation and monitoring systems used in underground mines by caving methods.

52

ICS-II, Task 4

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 dilution, 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 t he 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 2 2 17500 m , with an average of 3700 m .

(b)

As shown in Figure 9.2 the main causes of collapses are: o

(c)

Draw rate / Draw management

o

Structures

o

Mine planning / Mining sequence

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).

53

ICS-II, Task 4


1,5 m

CONCRETE DAMAGE

CONCRETE DAMAGE

Photograph 9.2:

Collapse at an extraction level drift of T eniente 4 Sur (1989).

0.40

0.35

0.30


0.25

0.20

0.15

0.10

0.05

0.00 0

2000

4000

6000

8000

10000

12000

14000

16000

18000

20000

AREA AFFECTED BY A SINGLE COLLAPSE (m 2 )

Figure 9.1:

Relative frequency of the area affected by a single collapse in underground mines by caving.

54

ICS-II, Task 4


1.00 0.95

FIRST MORE IMPORTANT CAUSE

0.90

SECOND MORE IMPORTANT CAUSE

0.85

THIRD MORE IMPORTANT CAUSE

0.80 0.75


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

WATER

MINE LAYOUT DESIGN

SUPPORT

MINE PLANNING MINING SEQUENCE

DRAW RATE DRAW MANAGEMENT

BLASTING

MA IN CA USE S OF A C OLL AP SE

Figure 9.2:

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

1.00

FIRST MOST COMMON REMEDIAL MEASURE

0.95

SECOND MOST COMMON REMEDIAL MEASURE THIRD MOST COMMON REMEDIAL MEASURE

0.90 0.85 0.80 0.75


0.70 0.65 0.60 0.55 0.50 0.45 0.40 0.35 0.30 0.25 0.20 0.15 0.10 0.05 0.00 IMPROVED GEOLOGICAL GEOTECHNICAL DATA

DRAINAGE

SUPPORT

MINE PLANNING MINING SEQUENCE


CONTROLLED BLASTING

MA IN R EM EDI AL M EA SU RE S FO R A CO LL AP SE

Figure 9.3:

Relative frequency of the different remedial measures against collapses that have been used in underground mines by caving.

55

ICS-II, Task 4


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 9.3:

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).

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).

56

ICS-II, Task 4


Photograph 9.5:

Example of HEAVY rockburst damage.

Photograph 9.5: Example of MODERATE rockburst damage.

57

ICS-II, Task 4


Photograph 9.7: Example rockburst damage.

of

LIGHT

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 presented below are based on the analysis of these data. The analysis of the data on rockbursts 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

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) m 200 (

T S R U B K C O R E L G N I S A Y B D E G A M A D H T G N E L T F I R D

HEAVY ROCKBURST DAMAGE TEN SUB 6 (1989 - 1992)

180

UNDERCUT LEVEL EXTRACTION LEVEL

160

VENTILATION LEVEL HAULAGE LEVEL

140

120

100

80

60

40

20

0

< 0

0 to 50

50 to 100

100 to 150

DIS TA NCE TO CA ING FR ONT (m)

Figure 9.4:

(d)

(e)

(f)

Relative frequency of the heavy rockburst damaged zones at different levels, and at different distances from the caving front.

As shown in Figure 9.5 the moderate damage at different levels varies with the distance to the caving front as follows: Level


Most Damaged Sector

Undercut

0 to 150 m

0 to 50 m

Extraction

< 0 to 150 m

100 to 150 m

Ventilation

< 0 to 150 m

< 0 to 150 m

Haulage

< 0 to 150 m

0 to 50 m

As shown in Figure 9.6 the light damage at different levels varies with the distance to the caving front as follows: Level


Most Damaged Sector

Undercut

0 to 150 m

100 to 150 m

Extraction

< 0 to 150 m

0 to 50 m

Ventilation

50 to 150 m

100 to 150 m

Haulage

< 0 to 150 m

0 to 50 m

As shown in Figure 9.7 the main causes of rockburst are: HIGH INFLUENCE:

Structures Stress environment Mining sequence Undercutting rate Draw rate

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) m 300 (


280 260 240 220

MODERATE ROCKBURST DAMAGE TEN SUB 6 (1989 - 1992) UNDERCUT LEVEL EXTRACTION LEVEL VENTILATION LEVEL HAULAGE LEVEL

200 180 160 140 120 100 80 60 40 20 0

< 0

0 to 50

50 to 100

100 to 150

DIS TA NCE TO CA ING FR ONT (m)

Figure 9.5:

Relative frequency of the moderate rockburst damaged zones at different levels, and at different distances from the caving front.

) m 400 (


380 360 340 320 300

LIGHT ROCKBURST DAMAGE TEN SUB 6 (19 89 - 1992) UNDERCUT LEVEL EXTRACTION LEVEL VENTILATION LEVEL HAULAGE LEVEL

280 260 240 220 200 180 160 140 120 100 80 60 40 20 0

< 0

0 to 50

50 to 100

100 to 150

D IS TAN CE TO C A IN G FR ON T (m)

Figure 9.6:

Relative frequency of the light rockburst damaged zones at different levels, and at different distances from the caving front.

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3

E C N A T R O P M I E V I T A L E R

2

1

0 STRUCTURES STRESS ENVIRONMENT

MINING SEQUENCE

UNDERCUTING RATE

DRAW RATE

ROCK MASS QUALITY

MINE LAYOUT

ROCK SUPPORT

UNDERCUTING MANAGEMENT BLASTING

MAIN CAUSES OF A ROCKBURST

Figure 9.7:

Relative importance of the different causes of rockbursts in underground mines by caving methods.

MODERATE INFLUENCE:

(g)

Rock mass quality Mine layout Rock support Undercutting management Blasting

As shown in Figure 9.8 the main remedial measures for rockburst are: HIGH INFLUENCE:

Monitoring Mining sequence Draw rate Contingency plans

MODERATE INFLUENCE:

Rock mass conditioning Draw management Support Technological improvements

9.3. SUBSIDENCE The connection of the cave back with the ground surface generates a subsidence crater, like the ones illustrated in Pictures 9.8 to 9.10, for sublevel, block, and panel caving mines. Usually the crater perimeter is subcircular, but in certain cases like the one shown in Picture 9.11, it could have a special shape. On the other hand the development of the subsidence crater could affect other mines located nearby, like the example shown in Picture 9.12. Of course, this will also be the case of any open pit mine developing a transition to underground mining by caving methods, like the example shown in Picture 9.13.

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3

E C N2 A T R O P M I E V I T A L 1 E R

0 MONITORING

MINING SEQUENCE

DRAW RATE

CONTINGENCY ROCK MASS DRAW PLANS CONDITIONING MANAGEMENT

SUPPORT

TECHNOLOGICAL IMPROVEMENTS

REMEDIAL MEASURES FOR ROCKBURSTS

Figure 9.8:

Photograph 9.8:

Relative importance of the different remedial measures for rockbursts in underground mines by caving methods.

Initiation of a subsidence crater due to the cave back connection to ground surface at a sublevel caving mine, Ridgeway, Australia.

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Photograph 9.10: Typical subsidence crater of a panel caving mine, Grasberg IOZ, Indonesia. Photograph 9.9:

Typical subsidence crater of a block caving mine, II Panel, Andina Mine, Chile.

Quebrada Teniente

Teniente 5 Pilares

Teniente 4 Fortuna

Teniente Sub 6

Teniente 4 Regimiento Regimiento

Teniente 3 Isla

Teniente 4 Sur

Photograph 9.11: Non typical, horse shoe shaped subsidence crater of a panel caving operation with several productive sector around a central pipe, El Teniente, Chile.

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OPEN PIT MINE DISPUTADA

SUBSIDENCE CRATER UG PANEL CAVING MINE III PANEL, ANDINA

Photograph 9.12: Example of the interaction between the subsidence crater of an underground panel caving mine (Andina), and a nearby open pit mine (Disputada), Chile.

PERIMETER OF THE SUBSIDENCE CRATER

Photograph 9.13: Example of the subsidence problem in a transition from open pit to underground mining by block caving, San Manuel, United States.

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DRIFT AFFECTED BY THE CRATER DRIFTS AFFECTED BYSUBSIDENCE THE SUBSIDENCE (INFLUENCE(INFLUENCE ZONE OFZONE) THE CRATER)

SURFACE SURFACE

CRATER CRATERWALL WALL

BROKEN ROCK BROKEN ROCK

z

Crater Crater Depth Depth

Crater Wall Mean Crater Wall’s Mean Inclination Inclination

Figure 9.9:

Parameters describing the overall geometry of a subsidence crater: crater wall depth, H, and crater wall mean inclination or break angle, α.

The overall geometry of a subsidence crater can be described as shown in Figure 9.9, by the depth (H) and mean inclination (α) of its walls will. The mean inclination of the crater wall is also known as break angle. The analysis of the data on subsidence indicates that: (a)

As shown in Figure 9.10, the height of the subsidence crater walls varies from less than 100 to 1400 m; but most data are below 700 m, and the average is about 450 (of course the height of a subsidence crater walls depends on the terrain topography).

(b)

As shown in Figure 9.11, the break angle defining the mean inclination of the crater walls varies from 40° to 90° (even in few cases there are overhanging walls), increased with rock mass quality, but most data are in the range from 50° to 90°, and if RMR > 70 the recorded break angles are all larger than 60°. It is very important to note that most of the data are for long term conditions (i.e. after the end of the block extraction).

(c)

Available data indicates that it could be possible to find a relationship between rock mass rating, RMR, the break angle, α, and the depth of the crater walls, H. This trend is illustrated in Figure 9.12. This topic will be included in the development of the geotechnical guidelines that are the second main activity of Task 4.

(d)

As shown in Figure 9.13 the main causes of subsidence are: HIGH INFLUENCE:

Structures Rock mass quality Block height Draw rate Draw management

MODERATE INFLUENCE:

Water conditions Footprint geometry Caving initiation Mining sequence Undercutting management

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0.20

0.15


0.10

0.05

0.00 0

100

200

300

400

500

600

700

800

900

1000

1100

1200

1300

1400

1500

DEPTH OF SUBSIDENCE CRATER WALL (m)

Figure 9.10: Relative frequency of different crater wall depths for underground mines by caving.

0.70 0.65 0.60 0.55


0.50 0.45 0.40 0.35 0.30 0.25 0.20 0.15 70 < RMR

0.10

60 < RMR < 70 50 < RMR < 60

0.05

40 < RMR < 50

0.00 0

10

20

30

40

BREAK ANGLE MEAN

50

60

70

80

90

ALUE (degrees)

Figure 9.11: Relative frequency of break angles for different rock mass qualities (Laubscher’s RMR), in the subsidence craters of underground mines by caving methods.

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1400

Laubscher's RMR 1300

41 to 50 51 to 60

1200

61 to 70

) 1100 s r e t e 1000 m (

H T P E D L L A W R E T A R C

> 71

900 800 700 600 500 400 300 200 100 0 40

45

50

55

60

65

70

75

80

85

90

B R EA K A NGL E (degrees)

Figure 9.12: Tentative relationship between the rock mass quality (Laubscher’s RMR), the break angle (α), and the depth of the c rater walls (H) in underground mines by caving.

3

E C N A T R O P M I E V I T A L E R

2

1

0 BLOCK HEIGHT

STRUCTURES ROCK MASS QUALITY

DRAW MANAGEMENT DRAW RATE

FOOTPRINT GEOMETRY WATER CONDITIONS

MINING SEQUENCE CAVING INITIATION

UNDERCUTING MANAGEMENT

M AIN C AU SE S O F S UB SID ENC E

Figure 9.13: Relative importance of the main causes of subsidence in underground mines by caving methods.

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3

E C N 2 A T R O P M I E V I T A L 1 E R

0 MONITORING

MINING SEQUENCE

DRAW RATE

RELOCATE BLASTING INFRASTRUCTURE

DRAW CONTINGENCY DRAINAGE MANAGEMENT PLANS

REMEDIAL MEASURES FOR SUBSIDENCE

Figure 9.14: Relative importance of the different remedial measures for subsidence in underground mines by caving methods.

(e)

As shown in Figure 9.14 the main remedial measures for subsidence are: HIGH INFLUENCE:

Monitoring Mining sequence Draw rate Relocate infrastructure

MODERATE INFLUENCE:

Blasting Draw management Contingency plans Drainage

9.4. WATER INFLOWS AND MUDRUSHES Water inflows and mudrushes have caused important damage to underground mines by caving methods, such as the example shown in Picture 9.14. These phenomena are sudden inflows of water and/or mud from drawpoints or other underground openings. Due to its own flow nature these phenomena propagate rapidly, endangering people, equipments, and infrastructure. Due to the fact that the amount of collected data on water inflows and mudrushes is limited, the analysis developed was mainly qualitative. The analysis of the data on water inflows and mudrushes indicates that:

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ICS-II, Task 4


Photograph 9.14: Example of the damage caused by a mudrush in an underground mine by caving methods.

(a)

(b)

As shown in Figure 9.15 the main causes of water inflows and mudrushes are: Most Frequent:

Wet ore / Clayey ore

Moderately Frequent:

Fine fragmentation Water collector crater Slope failures

Less Frequent:

Clayey overburden Warm rains / Rains Crown-Pillar failures Water above underground mine Backfill failures

As shown in Figure 9.16 the main remedial measures for water inflows and mudrushes are: Most Frequent:

Surface stabilization Draw management Contingency plans Drainage


Monitoring Technological improvements Relocate infrastructure Blasting

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0.25

0.20


0.15

0.10

0.05

0.00 CLAYEY ORE WET ORE

WATER-COLLECTOR CRATER FINE FRAGMENTATION

CLAYEY OVERBURDEN

SLOPE FAILURE

RAIN WARM RAIN

WATER ABOVE UG MINE CROWN-PILLAR FAILURE

BACKFILL FAILURE

MA IN C AU SE S OF MU DR US HE S A ND WA TE R IN FLO WS

Figure 9.15: Relative importance of the main causes of water inflows and mudrushes in underground mines by caving methods.

3

E C N 2 A T R O P M I E V I T A L 1 E R

0 DRAW MANAGEMENT CONTINGENCY SURFACE PLANS STABILIZATION

DRAINAGE MONITORING

TECHNO BLASTING IMPROVEMENTS RELOCATE INFRASTRUCTURE

REMEDIAL MEASURES FOR WATER INFLOWS & MUDRUSHES

Figure 9.16: Relative importance of the main remedial measures for water inflows and mudrushes in underground mines by caving methods.

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9.5. HANGUPS The ceasing of cave propagation due to the formation of stable geometry generates a hangup. This does not only stop the continuous breakage of rock, affecting the draw, but also generates a risk of a sudden failure of the stable geometry that could generate an air blast, causing important damage in the underground mine. After the failure of this stable geometry, usually the cave back reaches the ground surface generating a crater like the one shown in Picture 9.15.

Photograph 9.15: Example of the formation of a chimney crater immediately after the failure of the hangup at Inca West Sector, Salvador Mine, Chile (December 5, 1999).

The analysis of the data on hangups indicates that: (a)

As shown in Figure 9.17, the area of a hangup could vary widely, from less than 1000 2 2 2 m to more than 35000 m . Nevertheless, most reported data are below 15000 m , 2 and the average is 12000 m .

(b)

Reported data on air-blasts are related to hangups with areas larger than 10000 m .

(c)

As shown in Figure 9.18 the main causes of hangups are:

(d)

2

Most Frequent:

Geological changes Underestimation of rock mass quality


Low stress environment Changes in the undercut height

Less Frequent:

Draw rate / Draw management Undercutting sequence No measures taken to facilitate cave initiation Non standard undercut geometry

As shown in Figure 9.19 the main remedial measures for hangups are: Most Frequent:

Increase of the undercut area Conditioning the rock mass


Weakening of the boundaries of the hangup Draw rate / Draw management

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0.30

0.25


0.20

SALVADOR, SALVADOR, 05.12.1999 05.12.1999 0.15

0.10

0.05

AIRBLAST 0.00 0

5000

10000

15000

20000

25000

30000

35000

40000

HANGUP AREA (m2)

Figure 9.17: Relative frequency of different hangup areas in underground mines by caving.

0.35

0.30


0.25

SALVADOR, SALVADOR, 05.12.1999 05.12.1999

0.20

0.15

0.10

0.05

0.00 GEOLOGICAL CHANGES

UNDERESTIMATED ROCK MASS QUALITY

LOW STRESS ENVIRONMENT

CHANGES IN UNDERCUT HEIGHT


UNDERCUTTING SEQUENCE

NO MEASURES NON STANDARD TO FACILITATE UNDERCUT CAVING INITIATION GEOMETRY

MAIN C AUSES OF HA NGUP S

Figure 9.18: Relative importance of the main causes of hangups in underground mines by caving.

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0.45

0.40

0.35


0.30

SALVADOR, SALVADOR, 05.12.1999 05.12.1999

0.25

0.20

0.15

0.10

0.05

0.00

INCREASE UNDERCUT AREA

CONDITIONING THE ROCK MASS

WEAKENING THE BOUNDARIES


REMEDIAL MEASURES FOR HANGUPS

Figure 9.19: Relative importance of the main remedial measures for hangups in underground mines by caving.

9.6. FINAL COMMENTS The evaluation of the different causes for geotechnical hazards allows the grouping and classification of these causes as follows: Causes of Geotechnical Origin: Structures (presence of major geological structures) Rock mass quality (strength/cavability of the rock mass) Stress environment (in situ stresses) Water conditions (groundwater, hydrogeology, hydrology)    

Causes Related to Mine Design: Block height (column of solid rock above the UCL) Footprint geometry (area and geometry) Caving initiation (area, geometry, location, undercut height, measures to facilitate caving) Mine layout (drift orientation, spacing, size, draw point geometry and spacing) Support (drift support, draw point support, etc.)   

 

Causes Related to Mine Planning: Mining sequence (caving front orientation, geometry, advance direction, detentions, changes, etc.) 2 Undercutting rates (magnitude (m /month) and changes of magnitude) Draw rate (magnitude (m/day) and changes of magnitude, draw strategy) 

 

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Causes Related to Mine Operation: Undercutting management (tactical development of the undercutting: steps, operative sequence, short term rate) Draw management (tactical development of the extraction: uniformity, local changes, short term rate) Blasting (blast induced damage, remnant pillars, poor drilling/blasting operation) 





The evaluation of the relative frequency of each one of these causes in the reported geotechnical hazards, allow classifying them as causes of high influence, moderated influence, and low influence for each type of geotechnical hazard. This is shown in Figure 9.20, which indicates that the most important causes of geotechnical hazards are: STRUCTURES (geotechnical, present in all geotechnical hazards) DRAW RATE (mine planning, present in 4 of 5 geotechnical hazards) DRAW MANAGEMENT (mine operation, present in 3 of 5 geotechnical hazards) The evaluation of the different remedial measures for geotechnical hazards, allow the grouping and classification of these remedial measures as follows: Remedial Measures of Geotechnical Origin: Monitoring (seismic monitoring, displacements, field inspections, etc.) Rock mass conditioning (distressing, hydrofracturing, boundary weakening, etc.)  

Remedial Measures Related to Mine Design: Support (additional support, repair, backfilling, etc.) 

Remedial Measures Related to Mine Planning: Mining sequence (caving front orientation, geometry, advance direction, detentions, changes, etc.) Draw rate (magnitude (m/day) and changes of magnitude, draw strategy) Technological improvements (remote controlled equipment, new chute design, etc.) Surface stabilization (stabilization of slopes, mine waste deposits, etc.) Relocate infrastructure (changing the location of infrastructure affected by/causing a geotechnical hazard) 

   

Remedial Measures Related to Mine Operation: Draw management (tactical development of the extraction: uniformity, local changes, short term rate) Contingency plans (restrict access, evacuate people and equipment, sealing drifts or other cavities, etc.) Blasting (improved blasting techniques, blasting of remnant pillars, etc.) Drainage (surface and underground drainage measures, improving dewatering capacity, etc.) Undercutting management (tactical development of the undercutting: steps, operative sequence, short term rate) 



 



The evaluation of the relative frequency of each one of these remedial measures in the reported geotechnical hazards, allow classifying them as remedial measures of high influence, moderate influence, and low influence for each type of geotechnical hazard. This is shown in Figure 9.21, which indicates that the most important remedial measures for geotechnical hazards are: MINING SEQUENCE (mine planning, present in 4 of 5 geotechnical hazards) MONITORING (geotechnical, present in 3 of 5 geotechnical hazards) DRAW MANAGEMENT (mine operation, present in 3 of 5 geotechnical hazards)

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MAIN CAUSES

CAUSES

OF

GEOTECHNICAL

MINE DESIGN

PLANNING

OPERATION

GEOTECHNICAL HAZARDS IN R O C K

UNDERGROUND MINES BY CAVING S T R U C T U R E S

G E O T E C H N I C A L

M A S S Q U A L I T Y

S T R E S S

E N V I R O N M E N T

W A T E R

F O O T P R I N T

B L O C K

C O N D I T I O N S

H E I G H T

G E O M E T R Y

C A V I N G

I N I T I A T I O N

M I N I N G

R O C K

M I N E

S U P P O R T

L A Y O U T

S E Q U E N C E

U N D E R C U T I N G

R A T E

U N D E R C U T I N G

D R A W

R A T E

D R A W B L A S T I N G

M A N A G E M E N T

M A N A G E M E N T

COLLAPSES H A Z A R D S

ROCKBURSTS SUBSIDENCE WATER INFLOWS & MUDRUSHES HANGUPS

HIGH INFLUENCE

MODERATE INFLUENCE

LOW INFLUENCE

Figure 9.20: Relative influence of the main causes of geotechnical hazards in underground mining by caving.

MAIN

R E M E DI A L

REMEDIAL & MANAGEMENT

GEOTECH

M E A SU R E S

PLANNING

DESIGN

OPERATION

MEASURES

R

R

FOR

O

GEOTECHNICAL HAZARDS

C

S

K

UNDERGROUND MINES

M

BY CAVING M

A

O

S

N

S

I T

C

O

O

R

N

I

D

N

I

G

T

I C

I

H

A

A

R

N

N

C

T

A

O

E

E

W

G

A

I

S

I

M

W

M

T

N

A

P

A

F

N

B

R

A

P P

S

O

E

R

R

R

Q

A

O

I

A

G

T

U

T

V

L

S

E

E

E

E

I

T

M

M

Z

R

E

C

E

A

U

N

E

N

T

C

T

T

I

T

S

O

U

N

R

N

O N T I N

B

D

G

L

R

E

A

A

N

S

I

C

T

N

Y

I

A

N

G

G

E

P L A N S

E

COLLAPSES

E

N

D

U

I

H

C

R

N

C

O

F

N

G

E

C

R

C

S

N

T

L

E M

D

O

O

U

I

I

G

E

T

H A

ROCKBURSTS

Z A

SUBSIDENCE

R D S

WATER INFLOWS & MUDRUSHES

A L

HANGUPS

HIGH INFLUENCE

MODERATE INFLUENCE

LOW INFLUENCE

Figure 9.21: Relative influence of the main remedial measures for geotechnical hazards in underground mining by caving.

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10. CONCLUSIONS To develop this benchmarking 17 mines in 7 countries were visited, in order to collect relevant information. Also a comprehensive literature review was done to get complementary data, allowing to obtain information on up to 88 additional mines; nevertheless, in most cases the additional data do not include all the aspects considered in the benchmarking survey. 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. To make even easier the use of this information, Table 10.2 summarizes the current trends for the most relevant design parameters, and it can be used for pre-feasibility studies. It is also important to realize when a problem originates and when it is detected during the different stages of a mining project. As a first attempt, it is suggested use Table 10.1.

Table 10.1 POSSIBLE DISTRIBUTION OF THE ORIGIN AND DETECTION OF PROBLEMS DURING THE DIFFERENT STAGES OF A MINING PROJECT Stage of the Project

Origin of the Problem

Detection of the Problem

Pre-Feasibility Study

0% to 20%

0%

Feasibility Study

20% to 60%

0% to 5%

Basic Engineering

20% to 40%

0% to 5%

Detailed Engineering

10% to 20%

0% to 5%

Construction

10% to 20%

10% to 30%

Operation

20% to 60%

50% to 90%

(*) Modified from Sowers (1993)

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: 5.

CAVING PROPAGATION

6.

SUBSIDENCE

7.

CROWN-PILLAR

8.

WATER INFLOWS & MUDRUSHES

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Table 10.2 TYPICAL DESIGN PARAMETERS FOR A BLOCK/PANEL CAVING MINE Mine Design Parameter

Typical Value

Rock Mass Quality

50 ≤ RMR < 60

Acces

Decline

Block Height < 50000 m Footprint Area

50000 a 100000 m > 100000 m

Caving Initiation

l e v e L t u c r e d n U

2

This typical block height could vary ± 20%.

30000 m

2

75000 m

2

170000 m

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.

2

2

Area

10000 m

Shape

Square

Measures to Facilitate

Slot

Is highly recommended to facilitate cave initiation.

Hydraulic Radius

20 to 30 m

Avoid being close to the limit in Laubscher’s chart.

Spacing

15 m

Height Width

4m 4m


Undercut Height

8m

Could vary, but be careful if using small undercutting heights.

Undercut Rate

2100 m /month

s t f i r D

o l i t e c v a e r t L x E n

2

If RMR > 60 rock mass cavability must be evaluated carefully. Currently 70% of mines prefer declines, and 20% declines and shafts as mine access.

210 m 2

Comments

Smaller areas are not recommended, specially in massive rock masses. Internal corners must be avoided (e.g. a “L” shaped area).

This is the current practice.

2

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 Width

4m 4m

s t f i r D

Spacing

Draw Points

Influence Area Draw Rates

LHD Equipment


15 m 225 m

Could vary from 13 to 18 m. 2

2

Could vary from 169 to 324 m . 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.

0.20 m/day

Capacity

11 ton

It could vary ± 20%.

Traming Distance

140 m

Smaller tramming distances are preferable.

Powder Factor

400 grm/ton

Oversize Limit

3

Subsidence

RMR < 70 RMR > 70

1.8 to 2.0 m α >

45° α > 60°

For undercutting blasting. It could vary ± 20%. It could vary ± 20%. α 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 (see Figure 9.20 for main causes and Figure 9.21 for the most common remedial measures).

Instrumentation & Monitoring

The most common monitoring systems include displacements and seismicity. It is recommended 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 Laubscher’s 1990 system.

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11. ACKNOWLEDGMENTS The authors of this report want to thank all the sponsors for giving them the possibility of developing this work. The authors also want to express their sincere gratitude to all the colleagues that helped them to collect the information at each mine visited: Andina, Bingham Canyon, Cadia, Chuquicamata, El Teniente, Finsch, Grasberg Open Pit, Grasberg Underground (DOZ), Henderson, Kidd Creek, Kiruna, Koffiefontein, Mount Keith, Northparkes, Palabora, Ridgeway, Salvador. It is important to note that Grasberg, Henderson and Kidd Creek mines do not participate as sponsors of the ICS-II Study, but they were willing to help the authors in developing this benchmarking study. Finally, the authors want to acknowledge specially the support provided by Chuquicamata Mine of Codelco to develop this work.

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[118] Rojas, E.; Cavieres, P.; Dunlop, R. & Gaete, S. (2000): Control of Induced Seismicity at El Teniente Mine, Codelco-Chile, MASSMIN 2000, Proceedings, Brisbane, Qld, Australia. [119] Rojas, E.; Molina, R.; Bonani, A. and Constanzo, H. (2000):The Pre-Undercut Caving Method at the El Teniente Mine, Codelco Chile, MASSMIN 2000, Proceedings, Brisbane, Qld, Australia. [120] Rojas, E.; Molina, R. & Cavieres, P. (2001): Pre-undercut Caving in El Teniente Mine, Chile, Chapter 51 in UNDERGROUND MINING METHODS, Engineering Fundamentals and International Case Studies, Edited by Hustrulid and Bullock, SME. [121] Sandy, M. & Player, J. (1999): Reinforcement Design Investigations at Big Bell , ROCK SUPPORT AND REINFORCEMENT PRACTICE IN MINING, Villaescusa, Windsor & T hompson editors, A. Balkema. [122] Schroeder, K. (2002): Personal Communication. [123] Stawski, M. (2002): Personal Communication. [124] Stevens, C.; Sandbak, L. & Hunter, J. (1987): LHD Production and Design Modifications at the San th Manuel Mine, ROCK MECHANICS, Proceedings of 28 US Symposium on Rock Mechanics, edited by I. Farmer et al., Balkema. [125] Stewart, R. &d Sacrison, R. (1983): Geological Engineering – A Bridge Between Geologist and Miner , in Chapter 12 in SME - AIME Fall Meeting, Salt Lake City, Utah - October 19 - 21, 1983. [126] Streeton, G. (2000): Geotechnical Aspects of Open Stope Design at BHP Cannington, MASSMIN 2000, Proceedings, Brisbane, Qld, Australia. [127] Struthers, M.; Turner, M.; McNabb, K. & Jenkins, P. (2000): Rock Mechanics Design and Practice for Squeezing Ground and High Stress Conditions at Perseverance Mine, MASSMIN 2000, Proceedings, Brisbane, Qld, Australia. [128] Suarez, C. (2002): Personal communications. [129] Szwedzicki, T. (1989): Pillar Recovery at Manghura Copper Mines, Zimbabwe, Transactions of the Institution of Mining and Metallurgy, (Section A, Mining Industry), Volume 98, September-December 1989. [130] Tannant, D.; Martin, C. & Kaiser, P. (1997): Site Characterisation of the 6800 to 7800 Mining Block , Technical Report submitted to Falconbridge Limited - Kidd Mining Division by the Geomechanics Research Centre at Laurentian University. [131] Taylor, J. & Chinamatira, R. (2000): The Evolution of Sublevel Caving at Trojan Mine, Bindura, Zimbabw e, MASSMIN 2000, Proceedings, Brisbane, Qld, Australia. [132] Terbrugge, P. & Hanif, M. (1981): Discussion of a Large Scale on the Footwall of the Nchanga Open Pit, Zambia, Proceedings of the International Symposium on Weak Rock, Tokyo, September 1981. [133] Thompson, P. (1989): Slope Stability at Finsh Mine, ROCK SLOPE STABILITY, Sangorm Symposium, March 1989. [134] Tobie, R. & Julin, D. (1982): Block Caving: General Description, UNDERGROUND MINING METHODS HANDBOOK, edited by W. Hustrulid, SME. [135] Tobie, R.; Thomas, L. & Richards, H. (1998): San Manuel Mine, Chapter 43 in Techniques in UNDERGROUND MINING, Selections from Underground Mining Methods Handbook, Edited by R. Gertsh and R. Bullock, SME. [136] Turner, M. & Player, J. (2000): Seismicity at Big Bell Mine, MASSMIN 2000, Proceedings, Brisbane, Qld, Australia. [137] Uggalla, S. (2001): Sublevel Open Stoping – Design and Planing at the Olympic Dam Mine, UNDERGROUND MINING METHODS, Engineering Fundamentals and International Case Studies, Edited by W. Hustrulid and R. Bullock, SME. [138] Van As, A. (2002): Personal communication. [139] Van As, A. and Jeffrey, R. (2000): Caving Induced by Hydraulic Fracturing at Northparkes Mines, PACIFICS ROCKS, “ROCK AROUND THE RIM”, Proceedings of the Fourth North American Rock Mechanics Symposium: NARMS 2000, Washington; Girard, J; Liebman, M; Breeds, C and Doe, T . editors, A. Balkema. [140] Vera, S. (1981): Caving at Climax , Chapter 14 in Design and OPERATION OF CAVING AND SUBLEVEL STOPING MINES, Volume I, edited by D. Stewart, SME, New York. [141] Wilkes, T. (1992): The Support of Highly Fractured Rock in shallow Open Stopes at Koffiefountain Mine, Orange Free State, Republic of South Africa, ROCK SUPPORT IN MINING AND UNDERGROUND CONSTRUCTION, Kaiser & McCreath editors, Balkema.

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[142] Wilkie, N. and Holt, W. (1961): A Description of Open Pit Mining at Nchanga Consolidated Cooper th Mines Limited, Northern Rhodesia, Commonwealth Mining and Metallurgical Congress, 7 , May 1961, Johannesburg. [143] Wilson, A. (2000): The Past Focuses Support for the Future, MASSMIN 2000, Proceedings, Brisbane, Qld, Australia. [144] Wilson, D. (2002): Personal Communication. [145] Wood, P.; Jenkins, P. and Jones, I. (2000): Sublevel Cave Drop Down Strategy at Perseverance Mine, Leinster Nickel Operations, MASSMIN 2000, Proceedings, Brisbane, Qld, Australia. [146] Zappia, M. (1981): Gravity Caving and Production Hoisting at the San Manuel Mine, Chapter 16 in DESIGN AND OPERATION OF CAVING AND SUBLEVEL STOPING MINES, Volume I, edited by D. Stewart, SME, New York. [147] Zerga, R. (1981): Development and Production Equipment in Use at San Manuel, Chapter 50 in DESIGN AND OPERATION OF CAVING AND SUBLEVEL STOPING MINES, Volume II, edited by D. Stewart, SME, New York.

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