2

Question 1 (20 marks)

(a) Describe and illustrate the geotechnical classification method developed by Brady and Brown (1993)

for different underground mining methods - ranging from "naturally supported" at one extreme, to

"unsupported" at the other extreme. (4 Marks)

(b) What is meant by "naturally supported" and "unsupported", and what are the characteristics of each

of these categories in terms of orebody and near-field rock types, and stress states in the rock? (4

Marks)

(c) Illustrate this classification diagram by listing a number of representative mining methods (at least

two for each category) and explain why they are classified in this way. (4 Marks)

(d) What are some of the major core geotechnical risks associated with mining methods listed in (c)

above, at these two extreme ends of this classification system? (8 Marks)

3

Question 2 (20 marks) (a) A generic diagram of a slope in a jointed rock mass that is threatened by a planar slide is shown in the

sketch. The critical joint that will control the slide has a cohesive strength of 69 kPa and a friction

angle of 28°. The slope is located in moderate earthquake region and it is estimated that seismic

coefficient (g) of 0.15 is sufficient to account for potential earthquakes. Based on information

provided in Appendix 1:

(i) Compute the safety factor of slope according to the information provided in the figure

below. (6 Marks)

(ii) Will drainage make the slope stable? (2 Marks)

(iii) How much should the slope angle be decreased to stabilize it in drained condition to achieve

a safety factor of unity? (4 Marks)

(b) With the exception of drainage and flattening state any two strategies you may use to stabilize the

slope. (2 Marks) (c) State 6 sources of costs associated with slope failures in an open pit mine. (6 Marks)

156 m

45 32

Water table 10 m

123.32 m

4

Question 3 (20 marks)

(a) Explain the:

(i) aims for geotechnical engineering application and (2 Marks)

(ii) attributes of rock mass classification systems for geotechnical engineering application (2.5

Marks)

(b) A 5-m span by 5.2-m high drift is to be driven through granite at an underground mine site. The site

characterisation during exploration showed the granite has a dominant critical joint set dipping at 60°

against the direction of the excavation development.

Exploration borehole cores were tested as part of the characterisation of rock units encountered on

site for selection of the mining method and possibly whether to use drill and blast or mechanised

excavation. The site is remote and the company is concerned about cost. Hence, a decision was made

to use the point strength index (PSI) equipment instead of doing direct uniaxial compressive strength

testing. The point load testing was conducted on 36-mm diameter cores and gave a median Point

Load Strength Index of 8 MPa. An RQD of 70% was estimated in the direction of the haulage drift.

Assume water and stress and external factors and should not be accounted for in the Q system.

The onsite geologist is very meticulous and provided the following information in his logbook about

the structural characteristics of the granite:

Joint Spacing: 300mm

Number of joint sets 2 plus random

Joint persistence 4m

Joint Separation: 2mm

Joint Roughness: slightly rough

Joint Infilling: none (fresh)

Joint water condition: damp

(i) Determine the Rock Mass Rating (RMR) quality of the granite (2.5 Marks)

(ii) From the given information state what you need to determine the granite rockmass quality

using the tunneling quality index. (2 Marks)

(iii) Determine the tunneling quality index for the granite. (2 Marks)

(iv) Given that the tunnel is a permanent mine excavation, determine the excavation support

ratio. (2 Marks)

(v) Specify the support system to be used during tunneling in order to achieve a high advance

rate. (3.5 Marks)

(vi) Determine the final support system for the drift. (3.5 Marks)

(Note: See Appendix 2 and 3 for relevant information to assist you in providing the solutions)

5

Question 4 (20 marks)

Underground mining creates voids, which in most cases can be filled, and in many cases need to be filled.

Cement fill is used when complete ore extraction is planned and the stoping sequence primaries, secondaries

and in some cases tertiaries. Cemented fill is expensive and in some cases the cost may approach 25% of the

mining cost. The main cost component for the fill is the binder cost. The amount of binder needed at a

particular stoping environment depends on the required design strength of the fill. The design strength of the

fill is estimated by using appropriate methods of stability analysis. At present two types of method are used.

(a) List at least five different sources that can be used as backfill materials (2.5 marks) (b) State two methods used to estimate the design strength of cemented backfill (2 marks) (c) State the assumptions governing the Mitchel (1983) simplified limit equilibrium method for cemented

backfill stability in stopes (2 marks) (d) The figure below is a stope to be backfilled with cemented backfill. The geotechnical engineer needs

to estimate the uniaxial compressive strength of the backfill so that it does not cause dilution when the adjacent stopes are mined after it is cured. The bulk density of the cemented fill is estimated to be 24 kN/m3. Determine

(i) The height z (2.5 Marks) (ii) The required uniaxial compressive strength of the cemented fill (11 Marks)

DBH

DHBc b

244

2

Where c is cohesion, B is width of stope and H is the sublevel interval.

S

S

B=5 m

H=30 m

D=10 m

Z

45

6

Question 5 (20 marks)

(a) A proposal has been made to use an old underground limestone mine as a storage facility. It was mined

using the room and pillar method with a regular rectangular array of pillars. The clear spacing (side to side) between 7-m2 pillars is 6 m and the limestone is at the depth of 80 m below ground surface. The pillars were created using very careful blasting with no damage inflicted on them, giving a D value of one in Equations 1 to 3. D is defined as the rockmass disturbance factor which may be due to the excavation method used (e.g. poor or good blasting or mechanized).

The limestone is horizontally bedded with moderate spacing and gentle undulations. The bedding planes are smooth with slightly weathered surfaces and no visible aperture. Generally, the conditions inside the quarry are dry. According to the point load test of the pillar rock its uniaxial compressive strength is 100 MPa. The unit weight of the limestone is 28 kN/m3. (Note: GSI=RMR1989-5)

(i) Estimate the GSI for the pillar rockmass from RMR using Appendix 3 (2 marks) (ii) Determine the Hoek-Brown strength parameters mb and s for the rockmass using the

Equations (1) and (2) and assuming mi for limestone is 10: (3 Marks)

D

GSImm ib

1428

100exp

(1)

D

GSIs

39

100exp

(2)

3

20

15

6

1

2

1eea

GSI

(3)

(iii) Use mb and s from (ii) in Equation (4) (The Hoek and Brown failure criterion) to determine the pillar strength. (3 Marks)

a

ci

bci sm

3

31

(4)

(iv) Use the tributary area theory to estimate the average vertical stress in the pillar (2 Marks) (v) Using the results from (iii) and (iv) determine the factor of safety of the pillars (2 Marks)

(Provide any assumptions made in your analysis) (b) Describe what is meant by periodic weighting, associated with underground coal longwall mining operations. With the aid of sketches, discuss the:

(i) overall geotechnical mechanisms involved; (2 Marks) (ii) specific role of the longwall face supports in the mechanics of the problem; (2 points) (iii) range of conditions typically encountered; and (2 Marks) and (iv) various engineering, operational and design means available for dealing with the

problem. (2 points)

7

APPENDIX 1 – Equations planar slides in rock

𝐹𝑆 =𝑐𝐴 + [𝑤(𝑐𝑜𝑠𝛼 − 𝑎𝑠𝑖𝑛𝛼) − 𝑈 − 𝑉𝑠𝑖𝑛𝛼]𝑡𝑎𝑛𝜑

𝑤(𝑠𝑖𝑛𝛼 + 𝑎𝑐𝑜𝑠𝛼) + 𝑉𝑐𝑜𝑠𝛼

𝐴 = (𝐻 − 𝑧)𝑐𝑜𝑠𝑒𝑐𝛼

𝑤 = 0.5𝛾𝐻2 [(1 − (𝑧

𝐻)

2

) 𝑐𝑜𝑡𝛼 − 𝑐𝑜𝑡𝛽]

𝑈 = 0.5𝛾𝑤𝑧𝑤𝐴

𝑉 = 0.5𝛾𝑤𝑧𝑤2

APPENDIX 2 – Load Strength Index Is Correction to Is50

8

APPENDIX 3: Rockmass rating information RMR (Bieniawski, 1989)

A. Classification Ratings

Strength of intact rock

Point-load strength index (MPa)

* * * 1-2 2-4 4-10 >10

Uniaxial compressive strength (MPa)

<1 1-5 5-25 25-50 50-100 100-250 250

Rating 0 1 2 4 7 12 15

*For this low range uniaxial compressive test is preferred

Groundwater

Inflow per 10-m tunnel length (l/min)

>125 25-125 10-25 <10 none

Ratio 𝑗𝑜𝑖𝑛𝑡 𝑤𝑎𝑡𝑒𝑟 𝑝𝑟𝑒𝑠𝑠𝑢𝑟𝑒

𝑚𝑎𝑗𝑜𝑟 𝑝𝑟𝑖𝑛𝑐𝑖𝑝𝑎𝑙 𝑠𝑡𝑟𝑒𝑠𝑠 >0.5 0.2-0.5 0.1-0.2 <0.1 0

General conditions flowing dripping wet damp

Completely dry

Ratings 0 4 7 10 15

Drill core quality RQD (%)

Rating

<25 25-50 50-75 75-90 90-100

3 8 13 17 20

Spacing of discontinuities

Rating

<60 mm 60-200 mm 200-600 mm 0.6-2 m >2 m

5 8 10 15 20

Conditions of discontinuities

Rating

Use Table (B), or the following guidelines

Soft gouge >5 mm thick or separation

>5 mm, continuous

Slikensided surfaces or

gouge <5 mm thick or

separation 1-5 mm,

continuous

Slightly rough surfaces,

separation <1 mm, highly

weathered wall rock

Slightly rough surfaces,

separation <1 mm, slightly

weathered wall rock

Very rough surfaces, no separation,

unweathered wall rock, not continuous

0 10 20 25 30

B. Guidelines for classification of discontinuities:

Discontinuity length (persistence)

Rating

<1 m 1-3 m 3-10 m 10-20 m >20 m

6 4 2 1 0

9

Separation (aperture)

Rating

none <0.1 mm 0.1-1.0 mm 1-5 mm >5 mm

6 5 4 1 0

Roughness

Rating

very rough rough slightly rough smooth slickensides

6 5 3 1 0

Infilling (gouge)

Rating

Hard filling Soft filling

none <5 mm >5 mm <5 mm >5 mm

6 4 2 2 0

Weathering

Rating

unweathered Slightly weathered

Moderately weathered

Highly weathered

decomposed

6 5 3 1 0

C. Effect of discontinuity orientation in tunnelling:

Strike perpendicular to tunnel axis Drive with dip Drive against dip

Dip 45-90 Very favourable

Dip 20-45 favourable

Dip 45-90 fair

Dip 20-45 unfavourable

Strike parallel to tunnel axis Irrespective to strike

Dip 20-45 fair

Dip 45-90 very favourable

Dip 0-20 fair

D. Rating adjustment for discontinuity orientation:

Effect of discontinuity Orientation (from Table C)

Very favourable

favourable fair unfavourable Very

unfavourable

Ratings:

Tunnels and mines 0 -2 -5 -10 -12

Foundations 0 -2 -7 -15 -25

Slopes 0 -5 -25 -50 -60

E. Rock mass class determined from total ratings:

Rating 100-81 80-61 60-41 40-21 <20

Class no. I II III IV V

Description Very good

rock Good rock Fair rock Poor rock

Very poor rock

10

F. Interpretation of rock mass classes:

Class no. I II III IV V

Average stand-up time 20 yr for 15-

m span 1 yr for 10-m

span 1wk for 5-m

span 10h for 2.5-

m span 30min for 1-

m span

Cohesion of rock mass (kPa)

>400 300-400 200-300 100-200 <100

Friction angle of rock mass (deg)

>45 35-45 25-35 15-25 <15

11

APPENDIX 4 – Q-system Rating Q-system (Barton, 2002)

12

EXCAVATION SUIPPORT RATIO (ESR)

13

Chart fort selecting permanent support systems

14

Chart for selecting temporary support systems

‘END of EXAMINATION PAPER’