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D I G G I N G

D E E P E R DECEM

BER2002

1

Eden Paki is the Drill & Blast Supt at Newcrest

Mining Limiteds Cadia Valley Operations near

Orange in NSW where the current open pit

mining rate is around 80 Mtpa. Eden already

had substantial mining experience before moving

to Cadia, including 7 years D&B work for Roche

at KCGMs Superpit in Kalgoorlie followed by

working with Macmahon on D&B at the Mt

Todd Gold Mine in the Nor thern Territory.

Despite his more than full time job at Cadia,

Eden has also been acting as a sounding board

for Newcrests Telfer Project Feas Study open

pit mining team in the areas of assessing D&B

requirements for the project and how best to

PRACTICAL ADVICE WHERE ITS

CRUCIALLY NEEDED, AT THE FACECourtesy of one of AMCs Directors, Perth-based Lawrie Gillett, who cannot recommend

Eden Pakis work highly enough

satisfy those requirements. In the process of

helping the Telfer team, Eden has offered

guidelines based on his experience and these

are summarised below. They include formulae

which have been developed by others (apologies

in advance to authors of the formulae for not

being able to acknowledge them!!!) and these

formulae have gained good acceptance amongst

open pit mining D&B practitioners. Eden asks

that the reader bear in mind that the guidelines

were put together for practical, operations

based in-house assistance to his colleagues and,as such, do not pretend to be an exhaustive or

scientifically rigorous treatment of the subject.

Eden of course points out that there are always

site specifics which must be taken into account

in any mining determination, but AMC suggests

that Edens guidelines provide an excellent start

point. AMC has also found Edens guidelines

to be consistent with its discussions with drill

manufacturers and suppliers, blast hole drilling

contractors and drill bit suppliers.

We thank Eden for his cooperation in agreeing

to provide his guidelines to the broader mining

community through Digging Deeper.

Continued page 2

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As a base, I tend to favour the material size

required to allow for productive mining by the

selective loading toolsgiven that the hole

diameter determines the burden and spacing.

Until you can substantiate the mine to mill

process at the mine, then ultimately this should

not determine the drilling and blasting practices

(dont pour money into the process until you

can quantify and accurately measure it). Ground

hardness can also affect the hole diameter

selection process, but this can be addressed

through powder factors. Critical diameter for

explosives can also affect the hole diameter

but as long as you are looking at larger than

76mm blast holes, then this shouldnt be an issue.

Blast Size I always tend to look at going for

the largest achievable blast possible. This means

a reduction in infill drilling, reduction in initiation

costs, reduction in blasting frequency, which

means a reduction in blast affected production

delays. Altogether, a reduction in cost. When

talking blast size, I refer to the overall volume of

the blast. Vibration and timing (scheduling) and

physical restrictions such as the time taken to

load, tie-in and fire the blast play a more limiting

factor on the size blasts that I elect to use.

Blast Width I find this to be more important

than overall blast size. I have a tendency to look

at maximising the blast width, while minimising

the blast depth. Of course this is constrained

by the bench and mine plan, but where the real

estate is available to do so, I would elect to go

with 68 row shots that go on for as far as

possible. When looking at minimising the blast

depth, I refer to minimising to around 6 rows in

depthnot fewer. This provides excellent

displacement (if that is the desired result),

delivering excellent muckpile looseness.

Fragmentation is also maintained which means

that productivity is maximised.

2

EDENS RULES OF THUMB

There are a number of combinations that can

be used for Burden to Spacing ratios etc,

especially when blasting in unfamiliar terr itory.

When I encounter this, it is usually prudent to

go back to basics and apply a few well accepted

principles orrules of thumb. You will generally

find that most rules of thumb, require either

the hole diameter or burden to calculate what

other parameters will fit into the puzzle. The

following are fairly common rules of thumb

used throughout the industry when designing

blast patterns.

Burden = 25 to 35 hole diameters.

Spacing = 1.2 to 1.5 burdens (for largediameters e.g. Greater than 140mm) or around

1.5 to 1.8 for smaller diameters.

Stemming = 0.7 to 1.3 burdens. Keep in mind

that if you select a larger bench height, the

stemming column could have an impact on the

diggers productivity, especially if you are looking

at digging in flitches. For instance, if you were

to dig a 12m bench in three (3) flitches, then

the loading tool would be digging through the

stemming column. This is the area where you

are more likely to encounter oversize, thereforeproductivity would be much lower than the

other two (2) flitches.

Bench Height 4= 0 to 50 hole diameters.

Obviously this can usually change in accordance

with the bench height requirements of the

loading tools.

Subdrill = 8 hole diameters. This can be

increased for the face rows to around 10 to

12 hole diameters.

Powder Factors Governed by budget,

mine to mill processes and, to a lesser extent,

ground hardness.

Hole Diameters Essentially governed by the

required average material size. There are many

different perceptions on what determines the

hole diameter, but I personally select a hole

diameter based on the size of the loading

equipment, and any other material constraints

placed upon me by the milling process (finer

material means higher throughput). You must

find the middle ground between the two

at which point the higher cost of increased

powder factors, or drill & blast costs outweigh

the savings received from the milling process.

Blast Depth Blast depth can get you into

trouble, both through productivity issues and

through the cost required to minimise the effects

on productivity. I have trialed various blast depths

over time, trying various techniques such asincreased powder factors in the back rows, or

altered burdens and spacings, adding delays to

give a second front etc , however have had mixed

success in my effor ts. Essentially, the largest

impact of blast depth is muckpile looseness.

Higher powder factors can be introduced to

ensure that fragmentation is maintained in the

back rows of the blast, however it is very difficult

to provide a free face that allows sufficient

displacement to give productive dig rates.

Blast patterns that are greater than 8 rows,

are generally tighter to dig on the back few

rows. The lack of a free face in front, means that

the material is choked therefore becomes tighter

to dig. Additional heave is also a result which,

although providing some relief, is generally

confined to the upper reaches of the muckpile,

and not the lower and middle sections where

the shovel/excavator bucket spends most of its

time. The higher face created from the choked

rows also creates a hazard due to the higher face

height. Larger rocks, generally encountered in the

stemming collar are now higher in the diggingface, meaning that cooler packs on face shovels

become prone to accident damage.

Choke Firing Related to the above, choke firing

has its positives and its negatives. Choke firing

allows for minimal disruptions to the scheduling

process, as a digging unit does not need to free

face a blast and walk away. Instead, the digging

unit can continue digging through the area with

the choke shot being fired prior to the material

in front running out. The negative of choke firing

is muckpile displacement (which is actually apositive for grade control) and looseness

or lack of. Higher powder factors are generally

used also, to ensure that fragmentation is

maintained, as due to the confinement of the

blast, the energy tends to vent up and outwards.

This can mean poor fragmentation, especially at

the toe. Maximising free faces is always the best

option (productivity wise) however, choke firing

can be a necessary evilespecially in a gold

mining environment.

PRACTICAL ADVICE WHERE ITS CRUCIALLY NEEDED, AT THE FACEContinued from page 1

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Explosives Density As a general rule, the

harder the rock, the higher the explosives density.

You require more shock energy in harder ground.

For rock with a density of around 2.6 to 2.8, I

would look at using an explosive with a density of1.2 to 1.3. Further, as I am going to a high density

explosive, I would look to maximising the usage

of heavy ANFO based explosives as opposed

to pumped product, as the heave properties of

the augered product allows for better muckpile

looseness.

Number of blasts per blast day Determined

by labour. I would recommend blasting no more

than once per day, with no more that two (2) to

three (3) shots at a time. Any more and you start

to have problems relating to preparing the blaston the day (tie-ing in the shot), clearing the pit

for blasting, firing the shots and clearing the blasts

after firing. These can be very time consuming

and could result in a delay for blasting taking up

to an hour. I would not recommend planning to

tie in the blast the day before, purely for safety

reasons (also increases the risk of misfire through

human interaction or natural events such as wind

and storm).

Number of blasts per weekOnce again

I would look to minimise this. I try to scheduleblasting to a maximum of 3 times per week.

Any more and you start feeling the effects of

delays to production caused by blasting, plus the

strain starts to tell on the blast crew through

the frequency of tasks such as tie-ing in a shot.

This means that mistakes tend to creep in, plus

complacency becomes apparent. It is preferable

to dedicate whole days to purely charging and

stemming operations, which also gives the blast

crew adequate relief between each shot. Misfires

3

become a rare event, as opposed to a hazard

that is dealt with on a regular basis.

Stemming Size If you are looking at increasing

the blast hole diameter, I would recommend the

use of crushed rock for stemming. A rule of

thumb is that stemming size should be between

10% and 15% of the hole diameter. Drill cuttings

used in holes greater than 127mm tend to eject,

giving minimal confinement. The better

confinement offered by the crushed rock also

gives you the opportunity to reduce the

stemming column, thus improving the overall

fragmentation.

The table below is an example of the above rules

of thumb applied to an example set of project

parameters.

You will note that my Burden to Spacing ratio is

on the low sideI have actually found that

equilateral patterns give a great result (Burden =

0.867 of Spacing) however, they do cost a little

more in the long run. Please also note that I have

elected some fairly simple pattern sizes. Where

possible, I would suggest that you err away from

using dimensions such as 4.1 or 5.9. What tends

to happen is that operations will look for a

simpler dimension working in half metre

increments rather than 0.1 decimals.

Pulldown is another factor that you must

consider when selecting a hole diameter. This is

especially important when drilling rotary (note

that an 8 hammer is the largest commercial

hammer on the market todayother larger

hammers are generally made for specialist

purposes other than production blasthole

drilling). The pulldown requirement is usually

governed by the hole diameter, and the

compressive strength of the rock being drilled.

A basic formula to determine the pulldown

requirements of drilling with rotary is as follows:

P = (D x C) / 5

Where P = Required Pulldown, D = Hole

Diameter in inches and C = The rock uniaxial

compressive strength in PSI.

For example, if you were drilling a 251mm hole

in rock with a UCS of 260 MPa, the following

would apply:

251mm = 97/8; 260 MPa = 37,710 PSI

P = (97/8 x 37,710) / 5

P = 74,530 lbs

Using this number to assist you in drill selection

should go a long way to ensure that you have the

right drill combination to suit the application.

Imperative in this is selecting a drill that will

achieve your desired targets comfortably, unlike

the mistake made at one project where the drill

selection was aimed around maximising the

output of the drill through minimising the cost of

the initial purchase. Drill manufacturers will state

that their units are capable of drill hole diameters

ranging from 165mm to 270mm however, they

are not recommended to be run under load forextended periods of time. In order to achieve the

270mm, the engine, pumps and rotary head will

be under maximum load. The unit is capable of

achieving this, however you will find that over

time, in order to sustain this diameter, you are in

fact drastically reducing the life of the drill and its

components. Further, you may find that a dr ills

performance is reduced when operating for long

periods at the upper ranges of the drills

capabilities. I cannot provide operating data to

evidence this view, however it is a qualitative

observation made from experience.

The last thing to consider when looking at hole

diameters, is the size of the drill pipe associated

with that hole diameter. Drill pipe affects your

uphole velocity (also bailing or annular velocity)

which, in turn affects the performance of the

drilling consumables and the wear characteristics

of your drill string. A higher uphole velocity,

created by a smaller annulus will drastically

increase the wear on your hammers, subs and

drill rods, however a lower uphole velocity

may mean that youre not sufficiently lifting

the drill cuttings out of the blast hole .

Parameter 187 mm Diameter 251 mm Diameter

Pattern Burden x Spacing 5.5 x 6.5 6 x 7.5 7.5 x 9 9 x 10.5

Explosive Density (g/cc) 1.2 1.1 1.2 1.2

Hole Diameter (mm) 187 187 251 251

Burden (m) 5.4 6.0 7.5 9.0

Spacing (m) 6.5 7.5 9.0 10.5

Burden/Spacing Ratio 1.20 1.25 1.20 1.17

Bench Height (m) 12.0 12.0 12.0 12.0

Subdrill Length (m) 1.5 1.5 2.0 2.0Stemming Length (m) 4.5 4.5 4.5 4.5

Stemming/Diameter Ratio 24.1 24.1 17.9 17.9

Powder Factor (kg/bcm) 0.70 0.50 0.70 0.50

Explosive Mass (kg per hole) 297 272 564 564

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I have worked on 7000 to 8000 feet per minute

for fresh rock, going for the upper limits if that

material is also wet and down around 5000 feet

per minute for softer materials. A formula for

calculating uphole is as follows:

V = (Q x 183.33) / (D2 - d2)

Where V = bailing velocity, Q = compressor

output in CFM, D = Hole Diameter in inches and

d in drill pipe diameter in inches.

For example, if you are selecting a drill with a

compressor output of 1150cfm, drilling a 251mm

hole with 8 1/2 drill pipe, the bailing velocity

would be:

251mm = 97/8

V = (1150 x 183.33) / (97/82 - 81/22)

V = 8,345 ft/min

You can go to a smaller hole diameter to achieve

the same result, however you must ensure that

you have sufficient air capacity on the drill to

achieve an optimal bailing velocity, e.g. 1,450cfm.

Last but not least, when considering varying hole

diameters, consideration must be taken as to the

difference between the hole diameter and drill

pipe combinations available. The 81/2 drill pipe

used in the example above for the 251mm hole,

is too large to drill a 187mm hole. Therefore, in

order to use this configuration you would need

to change your drill string every time you wish to

change hole diameters. This requires labour,

cranes and, most important of all, time. The

optimal drill pipe size required to drill a 187mm

hole, will be too small to effectively run the

251mm rotary bits at the end of the daythis is

due to bailing velocity plus flex on the rods.

MESSAGE FROM

THE MANAGING

DIRECTOR

Redundancy

Redundancy is essential in engineering and

systems design. It is why a jumbo jet can lose an

entire electrical circuit and continue flying. It is

why your car brakes are still effective after losinga brake circuit.

When we prepare a feasibility study for a

new mine we specify a range of management,

technical and operating positions in an

organisation char t. Traditionally, the redundancy

allowance (sickness, leave and absenteeism)

was around 12% of the complement. This didn t

really allow for replacement of specialists, who

were required to arrange their work so that they

could take annual leave. We assumed that if

unusual circumstances led to higher absences,additional appointments would be made to

provide the necessary redundancy.

Most mines today operate on an organisation

chart that is pared to the bone. It shows just

enough people, if they were all available, to run

the mine safely and efficiently. But most mines

dont have those people available.

It has become common for a mine having a

technical and management complement of,

say, 15 to run chronically three or four positions

short. The cause is claimed to be outside

managements control, because there is a general

shortage of experienced professionals and they

are highly mobile. New vacancies arise as quickly

as positions are filled.

This phenomenon, in my opinion, compromises

the safe and effective working of the mine.

Production variability rises due to a lack of

planning and process control, leading to a loss

of profitability and to accidents. Both of these

outcomes should be unacceptable to

management and to boards of directors.

The solution is simple. Build some redundancy

into the management chart. If the bare bones

management chart requires 15 technical people,

make it your aim to employ 20. With the chronic

shortage still applying, you may sustain 15. The

same principle applies for operators. And guess

what? Safety and profitability will improve, despite

the higher employee cost.

I have been an expert witness in more than

thirty underground mine accident cases and

I believe that the cultural causes of accidents are

hard to quantify. However, it is easy to prove,

using employee records, that a mine was under-

staffed. I think that the causal link between

under-staffing and safety could be demonstrated.

Some operations are adequately staffed and

can demonstrate this. If your operation is not,

then I think that you are exposed. As a director

or manager you might turn your thoughts to

redundancy.

Merry Christmas.

Peter McCarthy

MINING AND TUNNELLING UNDER EXTREME CONDITIONS

mineABILITY will present a Mining and

Tunnelling under Extreme Conditions short

course in Melbourne 1415 April, 2003. The

course adopts an integrated approach to solving

problems encountered when mining and

tunnelling under extreme conditions. The five

presenters have a wide range of overseas and

local experience and will highlight the inter-

relationships between the rock mass character,

groundwater conditions, effective ground

stresses and feasible construction methods.

Case histories will be analysed that illustrate

the need for timely implementation of the most

appropriate ground remediation.

The course leaders are Dr Nick Barton from

Norway, who developed the widely used Q-

system for classifying rock masses and for

dimensioning rock tunnel and cavern support,

and Dr Eda Quadros from Brazil, who has over

25 years research experience in fluid flow

through jointed rock. Dr Barton has consulted on

numerous tunnel and cavern projects, reservoir

subsidence, rock stress measurement, nuclear

waste disposal and embankment dam projects

in a total of 26 countries over the last 30 years.

He has received several international awards for

his development work in rock tunnelling and

jointed rock behaviour.

For further details, contact David Pollard (phone

08 8362 5545, email dpollard@ausmin.com.au)

or Marnie Pascoe (phone 03 9670 8455, email

mpascoe@ausmin.com.au)

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Consider the geological aspects

Narrow reef/lode type structures are often too

narrow to require or retain any estimated grade

variation across the width of the zone. Therefore,

composites should span the entire width of the

zone with a single interval. It may be suitable to use

service variable methods for interpolation in order

to address the issue of variable sample support.

Where broader or mixed zones are being

modelled, the block dimensions become more

of a factor. It is rarely appropriate for thecomposites to be larger than the relevant block

dimensions, given the requirement for similar

support between block and composites.

However, composites can be smaller than

the block dimensions. In this case, the support

issues can be addressed through appropriate

discretisation parametersan array of points

within each block are used to generate individual

estimates which are then averaged to provide

the block estimate.

Compositing, through combination of data,inherentlysmooths the data and alters its

statistical characteristics. Sometimes compositing

is used to intentionally reduce the variance of

the data to reasonable levels, and hence improve

the quality of the variogram. The degree of

smoothing imposed by the compositing process

needs to be monitored. Before and after

compositing histograms of the grade data will

help to visualise the changes to the grade

distributions.

This is especially critical where the zone/domainbeing modelled is marginal to lower cut-offs likely

to be applied later on for repor ting.

Standard compositing routines tend to allow

specification of a minimum composite size to

retain, with some software defaulting to a value

that is half of the chosen composite length.

Where downhole compositing is used with

RESOURCE MODELLING:

DEVELOPING A STRATEGY FOR

COMPOSITING OF RAW SAMPLE

INTERVALS IN DRILLHOLE DATABy Ingvar Kirchner & John Tyrrell

One aspect of resource modelling that is

commonly misunderstood, or underestimated,

in terms of its significance, is the issue of

compositing of raw sample intervals in drillhole

data. Compositing is done specifically, to provide

common sample support for geostatistical

evaluations and grade interpolations.

Some of the issues that AMC routinely

considers when determining an appropriate

composite size are:

s The common raw sample lengths.

s Data quantity issues.

s Average zone/lode estimated true widths

and degree of internal definition required

from the estimates.

s Block size and bench heights (support issues).

s Data variability and the potential amount of

pre-estimation smoothing of the data.

s Loss of data at zone/domain margins.

s Selectively sampled/assayed intervals.

Generate a histogram of the raw

sample lengths in each zone/domain

Quite often the bulk of the sampling will occur

at some common interval (eg 0.5/1.0/2.0 metre

intervals etc). It would be unwise to use

a composite interval that splits up (or is smaller

than) the larger common raw sample intervals

termed decompositing. Decompositing of

larger intervals tends to cause a downward bias

in the data variance and causes some bias in the

rest of the statistical results. The initial effects can

be seen in variography (lower variance values,

increased ranges) and lower coefficient of

variations, and this can lead to other

downstream problems.

If the composite size is too large, the amount

of data available for the statistical analysis,

variography and grade interpolation will be

significantly reducedwith a corresponding drop

in the quality of the results.

composite runs controlled by zone/domain

boundaries, it is possible that critical data may be

lost at those boundaries. Narrow vein, lode, or

reef type zones may also inadvertently lose data

where the short length composite intervals fall

below the minimum composite length threshold

due to the narrow width of the zone. There are

residual retention type compositing processes

that can be used to overcome these issues

without resorting to very small minimum

composite lengths, which can produce a

negatively skewed distribution of sample lengths

and corresponding sample support problems.

A major issue is the effect of compositing on

selectively sampled or selectively assayed data.

Where data is incomplete and biased towards

higher grades (a common feature of the selective

sampling process), compositing has a nasty habit

of inflating those high grades and then causing

hot spots in the model, particularly where

those high grades are not bound by adjacent low

grades in the same hole. Dealing with selective

sampling/assaying issues is always hard, but

sometimes unavoidable. In this case, the best

strategy is probably to avoid using composite

intervals that are significantly larger than the

common sample intervals, thereby attempting

to minimise the inflation of those grades over

multiple/larger composite intervals.

It is apparent that different zones/domains may

require different compositing methods. Similar ly,

different elements may require different

compositing methods within the same model.

When choosing composite intervals and method,

a very considered, careful approach is required.

For further information, contact John

or Ingvar on jtyrrell@ausmin.com.au or

ikirchner@ausmin.com.au respectively

Comparison of Raw versus 2m Composites

Log Histogram 14/11/2002

SPEED UP YOUR SEARCH

AMCs website (www.ausmin.com.au) now has a

comprehensive consultant search facility which

provides a detailed and accurate listing of AMCs

consultants, who can be selected on the basis of

profession, expertise, commodity experience and

country experience (hyperlinks to email

addresses are provided). This new search facility

complements the full range of resumes available

on the website. Go to Staff on the sidebar and

click on Consultant Search

7/31/2019 dec02

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Australian Mining Consultants Pty Ltd ABN 58 008 129 164 Website: http://www.ausmin.com.au

MELBOURNELevel 19114 William StreetMelbourne 3000Telephone +61 3 9670 8455Facsimile +61 3 9670 8311amcmelb@ausmin.com.au

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UNITED KINGDOMHampton Business Centre7 Mount Mews, Hampton-on-ThamesMiddlesex, TW12 2SH, UKTelephone +44 20 8213 5881Facsimile +44 20 8979 2187amclondon@dial.pipex.com

In todays world the concept of 24/7 is

well established. Airlines talk about seat

utilisation factors, supermarkets are open

24 hours, and the local milk bar has been

replaced with a 7-Eleven, which never closes.

Why?

Its all about utilisation, the ability to use available

capacity fully. For airlines it means selling cheap

seats on the midnight flight, for supermarkets it

means dodging the shelf packers late at night,

and the 7-Eleven is well known for the midnight

munchies. So what does capacity utilisation mean

for mines?

In Optimisation we spend a lot of time talking

about capacity costs. Capacity costs are thosecosts that you incur when adding a unit of

capacity. Buy an additional loader and you have

to pay for manning, supervision, maintenance,

insurance, leases and ventilation. All these things

must be paid for, in full, before you can use the

loader. These upfront costs are the capacity costs

for the loader. That is they are the costs you have

to pay, to create the option of using the loader.

They are not as most people think, fixed costs.

If instead of adding a machine you take away a

machine, capacity costs become potential savings.

Fixed costs dont go away.

When we benchmark a mine one of the areas

we look at is production variability. We look at

two features, variability relative to budget, and

the variation on a month to month basis relative

to the peak monthly tonnage.

Figure 1 is the distribution of monthly tonnage

relative to peak monthly tonnage for the 25

underground mines on our database. The data

set ranges from Beaconsfield at 100 ktpa, to

Olympic Dam at 9 Mtpa.

The average is 83%. That is, on average,

underground mines use 83% of their

demonstrated capacity. Alternatively, in

Australian underground mines, there is on

average an opportunity to reduce capacitycosts by 15%. Our experience in benchmarking

is that demonstrated capacity is normally a lot

less than theoretical capacity.

The budget story is not much better. On

average, at the 3 standard deviation level,

mines deliver 20% of what they planned

to do. This is illustrated in Figure 2.

There are many causes for the level of variability.

However, they all boil down to a need to better

forecast actual production, and making sure that

alternative sources of ore are available to cover

for unexpected problems.

What does this mean to the consultant? Two

things. Firstly, when designing a mine and making

equipment selection, be aware of capacity

constraints. Try to avoid over sizing. Make sure in

schedules that multiple ore sources are available.

We normally find for the average mine (1.0 to 1.5

Mtpa), a minimum of 3 to 4 stopes are needed

to be available to bog at any one time. Less than

that, equipment tends to have lower utilisation.

The second is that when a client asks how they

can improve their performance, have a look at

their variability. It might not solve everything, but

it sure is a good start.

If you want to know a bit more about capacity

costs and variability, look up Johns paper entitledWhy Cost Cutting Fails to Deliver under Mine

Planning in AMCs Reference Library at

www.ausmin.com.au

Or for more information, contact

John de Vries, Principal Mining Engineer

with AMC, at jdevries@ausmin.com.au

GETTING WHAT YOU ARE PAYING FOR

A HALF FULL GLASS IS 50% UTILISATION

Figure 1: Production Relative to Peak Monthly Production 25 Underground

Mines 20002002Figure 2: Ratio of Actual to Planned Tonnage 25 Underground Mines 20002002