Conference PaperPDF Available

Design of stoping parameters and support system for longhole stoping method by numerical modelling

Authors:

Abstract and Figures

Longhole Stoping is a popular large scale underground mining method for metalliferous mine known for its ability to simplify stope sequencing. It encompasses advantage of Faster Stope Production Start leading to early return on investment. Ground support is a critical issue while designing of stoping parameters in this method. This can be achieved by analysis of stress conditions and formulation of guidelines for safe working. Under certain stress conditions, the use of pillars may be necessary when utilizing longhole stoping methods. In case of surface or sub-surface protection, properties of backfill to be used for filling voids are another important issue to be considered. Numerical modeling is a comprehensive approach to design stoping parameters including pillar dimensions, artificial supports and optimization of backfill materials. In the present paper authors have used 3-Dimensional numerical modelling technique to analyse the stoping parameters for Kayad underground mine of M/S Hindustan Zinc Limited. For estimating the stoping parameter rock mass properties of ore body and host rock have been tested in the CIMFR, laboratory. In situ stress measured by Messy India has been used as one of the basic input parameter. Proposed extraction mining proposition has been simulated for two lenses at shallower depth less than 150m. Based on the modelling results, support systems for drill level, draw levels and brow have been formulated.
Content may be subject to copyright.
DESIGN OF STOPING PARAMETERS AND SUPPORT
SYSTEM FOR LONGHOLE STOPING METHOD USING
NUMERICAL MODELLING TECHNIQUE
A. Kushwaha1, V K Himanshu2 and A. Sinha3
ABSTRACT
Longhole Stoping is a popular large scale underground mining method for metalliferous mine
known for its ability to simplify stope sequencing. It encompasses advantage of Faster Stope
Production Start leading to early return on investment. Ground support is a critical issue
while designing of stoping parameters in this method. This can be achieved by analysis of
stress conditions and formulation of guidelines for safe working. In case of surface or sub-
surface protection, properties of backfill to be used for filling voids are another important issue
to be considered. Numerical modeling is a comprehensive approach to design stoping parameters
including pillar dimensions, artificial supports and optimization of backfill materials.
In the present paper authors have used 3-Dimensional numerical modelling technique to analyse
the stoping parameters for Kayar underground mine of M/S Hindustan Zinc Limited (HZL).
For estimating the stoping parameter rock mass properties of ore body and host rock have been
tested in the CIMFR, laboratory. In situ stress measured by Messy India has been used as one
of the basic input parameter. Proposed extraction mining proposition has been simulated for
two lenses at shallower depth less than 150m. Based on the modelling results, support systems
for drill levels, draw levels and brow have been formulated.
1. Chief Scientist & Head, Bord & Pillar Mining Division, CSIR-CIMFR, Dhanbad
2. Student of Integrated M.Tech PHD & Trainee Scientist, AcSIR, CSIR-CIMFR, Dhanbad
3. Director, CSIR-CIMFR, Dhanbad
1.0 INTRODUCTION
Longhole stoping is a high production and low-cost
mining method for metalliferous mine. It is a very popular
method chosen when open pit mining activities are no
longer economical and mines move to underground
operations. Ground control is a critical issue while
designing of any stoping method and subsequent
systematic support have to be formulated for safe mining
practices. Numerical modeling is a comprehensive
approach for stability analysis of stope parameters based
on stress conditions of the adjoining zones of excavation.
Stress conditions are analysed based on physico-
mechanical properties of rock mass and in situ stresses.
At first modelled geometry are run for virgin condition and
then excavations are done in steps. Excavated part is
backfilled using Cemented rock fill and its physico-
5th ASIAN MINING CONGRESS
13 – 15 February 2014, Kolkata, India
The Mining Geological and Metallurgical Institute of India (MGMI)
2
Design of Stoping Parameters and Support System for Longhole Stoping Method Using Numerical Modelling Technique
mechanical properties are taken as input parameter.
Stability of stope parameter is assessed based on
application of rock failure criterion. Support parameters
are designed based on unstable zone found in the model
result and bearing capacity of rock/cable bolts.
Authors have used a case study for design of stoping
parameters and support system of Kayar underground
mine, HZL to extract two levels of K1A lens and Main lens
by underground stoping method. Drivages including Drill
levels and Draw levels will be driven in ore body itself for
the purpose of optimization of cost. In this purview, study
has been concentrated on stability of these drivages of
the proposed levels in both the lenses
2.0 LONGHOLE STOPING METHOD
This method is a variant of sublevel open stoping in
which longer blast holes with larger diameters (140 to 165
mm) are used. The holes are normally drilled using the
in-the-hole (ITH) technique. Depth of hole may reach up
to 100m. In this method, the miners must create a vertical
slot at one end of the stope and then work in the sublevels
to drill a radial pattern of drill holes. After a set of these
holes are loaded, blocks of ore body are blasted in to
open stope. There are two main variations of this method:
2.1 Longitudinal Longhole stoping
In Longitudinal Longhole stoping mining, stopes may
be retreated in the direction of the cross-cuts with a top-
down or bottom-up sequence. This is shown in figure 1.
In order to control dilution between individual stopes
along the strike, a top-down sequence normally requires
the installation of permanent rib pillars. A crown pillar also
needs to be used in order to control dilution and maintain
stability. This also means LHDs are operating under
backfill which is usually weaker than the rock. As well, it
is used to set apart any unconsolidated backfill that is
used in the upper stopes as the mining sequence
continues. For the bottom-up sequence, fill is required so
that there is a working floor as the extraction of the ore
continues upwards. If rib pillars are used along the strike
of the orebody along with backfill then a crown pillar is not
usually used. Two access crosscuts will cost more, but
increase the tonnage. The initial stopes may be located
in the centre of the mining block with retreats to follow
towards the abutments
Fig. 1: Sequence of stoping for longitudinal longhole stoping method
2.2 Transverse Longhole stoping
This is a bulk mining method in which the long axis
of the stope and access drifts are perpendicular to the
strike of the orebody. Draw points are located in the
undercut access drifts which extend from the footwall and
the free face is mined in a horizontal retreat from hanging
wall to the footwall. This method is used where the rock
mass quality of the hanging wall limits the length of the
open mining span. It has more flexibility with regard to
sequencing and scheduling due to presence of independent
access for each stope.
Stope sequencing in this method can be effectively
utilized to mitigate the effects of mining induced stress by
creating an active stress shadow, which can be managed
and manipulated to shelter existing and future excavations.
Most open stope mines practice transverse longhole
stoping, sequence their stopes based on the high stress
conditions of underground mining. One popular method
of mine sequencing known as the 1-5-9 sequence is
outlined in figure 2.
3.0 NUMERICAL MODELING APPROACH
Numerical modelling is a comprehensive approach
to study ground control stability and support design based
on instability occurred within the zone of extraction. There
are different Numerical simulation methods available for
this purpose. In the current study Finite Difference Method
using FLAC3D Software has been used. Design procedure
is followed based on Flow chart shown in Fig 3.
A. Kushwaha, V K Himanshu and A. Sinha 3
Fig. 2: Sequence of stoping for transverse longhole stoping method
4.0 CASE STUDY- KAYAR MINE, M/S HZL
Kayar mine is located at about 8-10 Km north-west
of Ajmer city in Rajsthan. Mine is operated by M/S
Hindustan Zinc Ltd. MECL drilled 42 holes (1994-97) on
promotional basis and estimated 10 million tonnes of
reserve grading 2.67% Pb and 14.89% Zn over a strike of
1 Km and vertical depth of 250 m. Hindustan Zinc Ltd
have proposed to extract these ores using underground
stoping method of mining.
There are two lenses – the Main lens and K1A lens.
The main host rock is Quartz mica schist with some
mineralisation also occurring in calc silicate. Dip of the
ore body is around 75º towards east. Main lens has been
dissected at many places by pegmatite. Description of
these two lenses is given in Table 1. Figure 4 shows
longitudinal vertical section of these ore lenses with
proposed stoping plan. In the first instance, mine
management have proposed first trial extraction of K1A
lens from 363 mRL to 379.5 mRL and then Main lens from
325 mRL to 375 mRL.
Table 1: Description of ore lenses present at Kayar Mine
Strike length (m) Av. width (m) Depth from (m) Depth up to (m)
Main Lens Steeper portion 900 5 450 230
Shallower portion 800 20 230 50
K1A 250 4 470 350
4.1 Trial Stoping Sequence for K1A Lens
Stoping sequence for K1A lens is proposed as shown
in Figure 5. One Drill level and one draw level of height
4.5m and width 5m each is driven. Strike length of the
stope is 75m. This 75m Stope will be divided in to two
stopes of 37.5m strike length each and maximum width
of extraction will be 7m. Draw point will be at the junction
of two stopes. Draw point will be connected through a
cross cut from where material transportation will take
place. Stope 1 will be extracted first and after complete
extraction it will be backfilled with Rock Fill (RF) &
Cemented Rock Fill (CRF) combination, then stope 2 will
be extracted and backfilled.
Fig. 3. Flow chart showing procedure of support design by Numerical
Modeling Technique.
4
Design of Stoping Parameters and Support System for Longhole Stoping Method Using Numerical Modelling Technique
4.2 Stoping Sequence for Main Lens
Stoping sequence for Main lens is proposed as shown
in Figure 6 .One Drill level and two draw levels of height
4.5m and width 5m each is driven. Length of the each
stope in the strike direction has been taken as 25m with
maximum width of extraction 7m. Extraction sequence
has been proposed from two ends in south and north side
from S13 both separately as shown in Fig. 4. There will
be four draw points at the junction of different adjacent
stopes. Draw points will be connected through a cross cut
from where material transportation will take place. Stope 1
will be extracted first and after complete extraction it will
be backfilled, then stope 2 will be extracted and backfilled
and so on. First Draw level will serve as Draw level for
Stope1, Stope2 & Stope3 and Second Draw level will serve
as Drill level for these three stopes. Second Draw level
will serve as draw level for Stopes 4, 5 & 6.
4.3 Scope for Stability analysis of Stoping
Parameters
(1) Stability analysis and support design for Drill level
and Draw level of K1A Lens and Stope brows during
extraction of K1A lens stopes.
(2) Stability analysis and support design for First draw
level and Second draw level during extraction of
stopes 1, 2 & 3 of Main Lens and Drill level during
extraction of stopes 4, 5 & 6 of Main Lens.
(3) Stability analysis and support design for Stope brows
during extraction of Main lens stopes.
4.4 Rock mass properties of mine
Mine management has collected core samples from
two boreholes such as KGT1 and KGT2 drilled in the
footwall of the ore body and sent to CIMFR Dhanbad for
testing in the laboratory as summarised in Table 2.These
physico-mechanical properties have been used for
modeling purpose.
Table 2: Physico-mechanical properties of
host rock
Properties Host Rock
RMR 65
Compressive Strength óc50 MPa
Tensile Strength ót5 MPa
Young’ modulus, E 10 GPa
Poisson’s ratio 0.22
Rock density 2750 Kg/m3
Excavated stopes have been backfilled with cemented
rock fill with the properties are given in Table 3.
Table 3: Physico-mechanical properties of
cemented rock fill
Young’s Modulus, MPa Poisson’s ratio Density, Kg/m3
194 0.2 2000
4.5 In situ stresses
In situ stress values have been taken from report of
Hydro-fracturing test done at HF2 by M/S Messy India
Ltd. supplied by the mine management.
Fig. 4: Longitudinal Vertical section of Kayar Deposit along with
proposed stoping plan.
Fig.5: Stoping sequence for K1A Lens.
Fig. 6: Stoping sequence for Main Lens.
A. Kushwaha, V K Himanshu and A. Sinha 5
Horizontal in-situ stresses
SH = 10.59 + 0.0544*(H-67) MPa
Sh = 5.35 + 0.0238*(H-67) MPa
Vertical in-situ stress
Sv = 0.0275 H MPa
Where SH and S-h are the major and minor horizontal
stresses (MPa), Sv vertical stress (MPa) and H is the depth
of cover (m). The direction of major horizontal stress is
almost along the dip direction.
4.6 Numerical modelling
Numerical modelling has been conducted using
Finite Difference method. A Finite Difference Software,
viz., FLAC3D of ITASCA Consulting Groups Inc.,
Minnesota, USA have been used for this purpose. The
safety factor contours are evaluated using a programming
environment (FISH) of the above software taking CIMFR
failure criterion in to account.
4.6.1Modelling Results for K1A Lens
3-D modelling has been done for the extraction of
12m parting between Drill level and Draw levels. Model
geometry is made on the basis of stoping pattern shown
in figure 5 taking width of the stoping as 7m. For the
purpose of long term stability of Drill Level and Draw Level
safety factor of 1.5 is to be maintained. Figure 8a & 8b
shows contour of safety factor in and around the both
drivages while extraction and subsequent filling of stope1
and extraction of stope 2 upto 20 m. From the figure 8b, it
is clear that safety factor less than 1.5 is limited up to 1.5
m height from drill level, similarly from figure 8a, it is clear
that safety factor less than 1.5 is limited up to 1.5m height
from draw level. Stability of stope brow is necessary for
short term, so safety factor 1.0 is sufficient. From figure
8a, safety factor of 1.0 is limited up to height of 1.0 m from
stope brow.
4.6.2Modelling and Results of Main Lens
3D modelling has been done for the extraction of
75m stope between 325mRL to 375mRL Model geometry
is made on the basis of stoping pattern shown in figure 6
and as explained in section 4.2.2.
Figure 9a & 9b shows contour of safety factor while
extraction and subsequent filling of stope1 & stope 2 and
extraction of stope 3 up to 10 m. From the figure 9b, it is
clear that safety factor less than 1.5 is limited up to 2 m
height from second draw level; similarly from figure 9a, it
is clear that safety factor less than 1.5 is limited up to 2m
height from first draw level. Stability of stope brow is
necessary for short term, so safety factor 1.0 is sufficient.
From figure 9a safety factor of 1.0 is limited up to height
of 2 m from stope brow.
Figure 10a & 10b shows contour of safety factor
while extraction and subsequent filling of stope1, stope 2,
stope 3, stope 4, stope 5 and extraction of stope 6 upto 10
m. From the figure 10b, it is clear that safety factor less
than 1.5 is limited up to 1.8 m height from drill level,
similarly from figure 10a, it is clear that safety factor below
1.5 is limited up to 2m height from second draw level.
From figure10a, safety factor of 1.0 is limited up to height
of 2 m from stope brow.
Fig. 8a: Isometric View of Contour of safety factor for backfilled stope
1 and partially extracted stope2 of K1A Lens
Fig. 8b: Side View of Contour of safety factor for backfilled stope 1
and partially extracted stope2 of K1A Lens
6
Design of Stoping Parameters and Support System for Longhole Stoping Method Using Numerical Modelling Technique
4.9 Design of Support system
4.9.1Design of support system for Drill Level &
Draw Level of K1A Lens
Block contour of safety factor (see fig 8a & fig 8b)
over drill level at 370.5mRL and draw level at 363mRL for
stope dimension of 7m wide and 12m height indicates
factor of safety 1.5 confined within 1.5m of the drill level.
So rock load density can be computed as RL = ñ x h1.5,
where density of rock ñ is 2.75 t/m3.
RL = 2.75 x 1.5 = 4.125 t/m2
It is proposed to support the Drill level with 2.4m
long fully resin encapsulated rock bolts having bearing
capacity (bc) of 20 tonne. In view of stability required for
the levels, it was suggested that 4 number of full resin
encapsulated rock bolts(n) of length 2.4m should be
grouted at a spacing(Sp) of 1.33 m between two
consecutive rows in 5m wide span of the stope. Hence
Applied support load density (ALS) in the stope back can
be computed as
ALS = (20 x 4)/ (5 x 1.33) = 12.03 t/m2
So safety factor of each bolt will be 12.03/4.125 =
2.91>1.5 required for long term stability. Plan and sectional
view of proposed pattern of bolting is shown in Figure 11.
Figure 9a: Isometric View of Contour of safety factor for backfilled
stopes 1& 2 and partially extracted stope3 of Main Lens
Fig. 9b: Side view of Contour of SF for backfilled stopes 1& 2 and
partially extracted stope3 of Main Lens
Fig. 10a: Isometric view of Contour of SF for backfilled stopes 1, 2, 3,
4, & 5 and partially extracted stope 6 of Main Lens
Fig. 10b: Side view of Contour of SF for backfilled stopes 1, 2, 3, 4, &
5 and partially extracted stope6 of Main Lens Fig. 11: Plan and sectional view of support system for Drill Level &
Draw Level of K1A Lens
A. Kushwaha, V K Himanshu and A. Sinha 7
4.9.2Design of support system for Stope Brow of
K1A Lens
Block contour of safety factor (see fig 8a) over stope
brow indicates factor of safety 1.0 confined within 1.0m
from it. So rock load density can be computed as RL = ñ x
h1.0, where density of rock ñ is 2.75 t/m3.
RL = 2.75 x 1.0 = 2.75 t/m2
It is proposed to support the stope brow with 2.4m
long fully resin encapsulated rock bolts having bearing
capacity (bc) of 20 tonne. In view of short term stability
required for the brow, it is suggested that 3 number of full
resin encapsulated rock bolts(n) of length 2.4m should
be grouted at a spacing(Sp) of 2.0 m between two
consecutive rows in 5m wide span of the stope. Hence
Applied support load density (ALS) in the stope back can
be computed as
ALS = (20 x 3)/ (5 x 2.0) = 6 t/m2
So safety factor of each bolt will be 6.0/4.125 = 2.18
>1.0 required for short term stability. Plan and sectional
view of proposed pattern of bolting is shown in Fig 12.
4.9.3Design of support system for Drill Level &
Draw Levels of Main Lens
Block contour of safety factor over drill level at 375
mRL and draw levels at 329.5 mRL and 352.5 mRL for
stope dimension of 7m wide and 50m height indicates
factor of safety 1.5 confined within 1.8m for drill level and
2m for other cases (fig 9a, 9b, 10a & 10b). Safety factor
contours shows similar results for both the lenses due to
presence of similar stope width and Drivages width and
working is at the similar depth. So support system design
for draw levels will be similar as for support system of
K1A Lens and will be as per Fig.11.
Rock load density for drill level can be computed as
RL = ñ x h1.5, where density of rock ñ is 2.75 t/m3.
RL = 2.75 x 1.8 = 4.95 t/m2
It is proposed to support the stope brow with 2.4m
long fully resin encapsulated rock bolts having bearing
capacity (bc) of 20 tonne. In view of stability required for
the levels, it is suggested that 4 number of full resin
encapsulated rock bolts(n) of length 2.4m should be
grouted at a spacing(Sp) of 1.33 m between two
consecutive rows in 5m wide span of the stope. Hence
Applied support load density (ALS) in the stope back can
be computed as
ALS = (20 x 4)/ (5 x 1.33) = 12.03 t/m2
So safety factor of each bolt will be 12.03/4.95 =
2.43>1.5 required for long term stability. Plan and sectional
view of proposed pattern of bolting is shown in Fig 11.
4.9.4Design of support system for Stope Brow of Main
Lens
Block contour of safety factor (see fig 8) over stope
brow indicates factor of safety 1.0 confined within 2.0m
from it. So rock load density can be computed as RL = ñ x
h1.0, where density of rock ñ is 2.75 t/m3.
RL = 2.75 x 2.0 = 5.5 t/m2
It is proposed to support the stope brow with 2.4m
long fully resin encapsulated rock bolts having bearing
capacity (bc) of 20 tonne. In view of short term stability
required for the brow, it is suggested that 3 number of full
resin encapsulated rock bolts(n) of length 2.4m should
Fig. 12: Plan and sectional view of support system for Stope Brow of
K1A Lens & Main Lens.
8
Design of Stoping Parameters and Support System for Longhole Stoping Method Using Numerical Modelling Technique
be grouted at a spacing(Sp) of 2.0 m between two
consecutive rows in 5m wide span of the stope. Hence
Applied support load density (ALS) in the stope back can
be computed as
ALS = (20 x 3)/ (5 x 2.0) = 6 t/m2
So safety factor of each bolt will be 6.0/5.5 = 1.09
>1.0 required for short term stability. Plan and sectional
view of proposed pattern of bolting is shown in Fig 12.
5.0 CONCLUSIONS
Longhole stoping method is a popular method for
high production and low cost investment. Ground control
is a critical issue while making any stoping method
operational. Numerical modeling is a comprehensive
approach to assess stability parameters for stoping and
formulation of systematic support rule depending on stress
behaviour. Safety factor analysis and study of stability
behaviour of surrounding rocks at Kayar mine using
Numerical modeling recommends supporting of Drivages
and stope brow as per pattern discussed in study.
Continuous monitoring of strata is also recommended.
6.0 ACKNOWLEDGEMENTS
The authors would also like to thank Mr. Rana
Bhattacharjee & Mr. Subhashish Tewari, Technical Officer,
Bord and Pillar Mining Division, CIMFR for their valuable
support. Authors would also like to thank Mine
Management, Kayar Mine of M/S Hindustan Zinc Limited
for providing necessary data required for study.
7.0 REFERENCES
1. Howard L. Hartman, Introductory Mining
Engineering, pg no.- 366 to 372
2. Willium A. Hustrulid, Richard L. Bullock,
Underground Mining Methods, pg no.- 7 to 8
3. https://queensminedesign.miningexcellence.ca/
index.php/Transverse_longhole_stoping
4. https://queensminedesign.miningexcellence.ca/
index.php/Longitudinal_longhole_retreat
5. Manual of FLAC 3D Software, supplied by Itasca
Consultants, USA
6. CIMFR testing report of Kayar Mine rock.
7. Geological Report of the Kayar Mine supplied by
Kayar Mine Management.
8. In situ stress measured by Messy India, Provided
by Kayar Mine Management.
9. Sheorey P. R (1997), Empirical rock failure criteria,
A.A. Balkema Rotterdam, 176p.
Conference Paper
Full-text available
Longhole stoping with ring drilling is best suited method for large scale production from exploitation of nearly vertical ore deposit. However production blasting in underground stope beneath surface structure is a challenge for blast designers. Residential and industrial surface structures above underground stopes experiences substantial cracking due to high magnitude of blast vibration. Directorate General of Mines safety (DGMS), India has framed regulation to limit blast vibration upto stipulated level to safeguard these surface structures. Following paper deals with designing of controlled blasting parameters for blasting at underground long hole stope with safety of surface industrial structures. The experimentation with blast design has been performed for this purpose at Sindesar Khurd Mine of M/s Hindustan Zinc Limited, Rajsthan, India. Experimental and simulation approach has been used to optimize blast design parameters to safeguard surface industrial structures like water tank, electrical substations, crushers and milling plant. This has been achieved by optimization of charging parameters and delay sequence. Simulation approach predicted blast induced vibration of 26mm/s at a distance of 100m from simulated signature blast hole. The predicted vibration is from model simulated considering equivalent rock mass as per mine condition. However, Anisotropy and heterogeneity of rock mass tend to decrease blast vibration around structure. Charging parameters have been designed by experimental blasts. Altogether 64 experimental blasts were conducted with varying blast design parameters. Charge weight per delay has been optimized using USBM predictor equation. Multivariate statistical analysis approach has been used for optimization of total explosive charge in a blasting round. Multivariate predictor equation has been developed considering hole diameter, number of blast holes in a blasting round, total explosive charge in a blasting round and distance of structure from blast face as independent variables and peak particle velocity of ground vibration as dependent variable. Total explosive charge and explosive charge weight per delay has been tabulated based on these predictors. Results have been recommended for day to day blasting at mine in order to reduce blast vibration near surface structures. Delay timing between blast holes has been optimized by near field blast vibration monitoring. Analysis of waveforms for recorded near filed blast vibration data reveals that 40ms is the optimum delay for ring blasting in order to reduce vibration. Blast design has been recommended for stoping based on the results of simulation and statistical analysis.
Conference Paper
Full-text available
Purpose of the study: To Design safe stoping dimension for extraction of ore body in underground metaliferrous mine. Principal results: Stoping of ore body causes failure around excavation zone. Assessment of failure around wall rocks, stope brow and stope back gives idea for design of stoping dimension in strike and dip direction prior to backfilling. Major conclusions: Stoping is final extraction of orebody from an underground mine. This operation leads to redistribution of stresses around stoped out area. Significant deformations in wall rock, stope back and stope brow is observed due to stoping. These deformations can be taken care of by supporting the structures or backfilling voids generated due to stoping. Stoping dimensions designed here include extension of stopes in strike direction (stope length) and its extractionbackfill sequence. Following paper deals with a case study for design of stope dimension of an underground mine based on Mohr coulomb elasto-plastic failure analysis using numerical modeling method. Numerical models have been simulated for different extraction and backfill sequence for this purpose and failure around wall rock, stope back and stope brow have been analysed. Recommendations have been made for safe extraction of orebody based on observed data.
Article
Provides treatment of the applications of mining engineering while reinforcing material with analyses of special topics as well as numerical examples and problems. Initial chapters are devoted to fundamentals, explaining the four stages of mining - prospecting, exploration, development, exploitation - and the unit operations of mining. The text continues with coverage of surface mining and underground mining.