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COPPER HEAP LEACH TESTING, INTERPRETATION AND SCALE UP

Graeme Miller and Tim Newton

ALTA Copper Hydrometallurgy Forum, QLD, 1999.

CONTENTS

1. ABSTRACT 2

2. INTRODUCTION 3

3. TOTAL COPPER VS RECOVERABLE COPPER 4

4. TEST PROGRAMME OVERVIEW 5

5. COLUMN TEST DESIGN 9

6. PILOT LEACH HEAPS 13

7. CALCULATION AND PRESENTATION OF RESULTS 14

8. INTERPRETATION OF RESULTS 16

9. SCALE UP 19

10. CONCLUSIONS 21

11. ACKNOWLEDGEMENTS 22

12. NOMENCLATURE 23

13. REFERENCES 24

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1. ABSTRACT

The testing of copper heap leach systems has undergone radical revision in the last five years

in an effort to obtain better interpretation and modelling of field results. A number of workers

have stressed the importance of establishing a project specific data base of different

characterisation and leaching responses which enables proper planning of the mining and

placement sequence.

Testing methods have been refined to allow information to be gathered on both the early

leaching and that carried out in mature leach solutions with high concentrations of dissolved

gangue mineral ions. Both are important in establishing the leaching rate and the acid

consumption rate. Further testing methods have been developed to more accurately define the

different leaching regimes of oxide and sulphide dominated minerals. Use of appropriate

bacterial cultures is needed to properly leach the secondary sulphides.

Interpretation of the leaching tests has been greatly assisted recently by the availability of

excellent data on large-scale field leaching trails from a number of Australian copper operations.

This has allowed the development of two distinct models that allow the oxide and sulphide

mineral leaching to be interpreted more reasonably. The method of interpretation takes into

account not only the copper extraction kinetics but also the acid consumption dynamics.

Sufficient test work needs to be planned to ensure that the acid consumption response to

alterations in acid concentration can be assessed.

The scale up from columns to field expectations still relies on the application of some factors to

account for the non-ideal conditions encountered in the field. Some previous work has been

reassessed to provide an idea of the effect of this on the leaching parameters. The copper

extraction and acid consumption models have been used as the basis of scale up to predict the

field response to the design conditions. With sufficient test work knowledge the effects of

altering the acid concentration during the leach cycle can be assessed and applied to the

economic modelling of the project. These new tools also allow the full assessment of operating

parameters such as heap height, head grade, material bulk density and irrigation rate before

decisions on the selected values are finalised.

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2. INTRODUCTION

To ensure the success of copper heap leaching projects, the metallurgical characteristics of the

orebody must be well defined. An extensive test programme should be initiated in the early

stages of development to determine the effect on the project’s feasibility, and to provide

information needed for the plant design. Where a bankable feasibility document is required, the

results must stand up to the scrutiny of third parties.

Orebody/reserve data, basic mineralogy and lithology should be known prior to planning the

leach test program. Representative samples are then required for each ore category. The

testing program can then proceed in stages, starting with more detailed mineralogy, size

distribution, elemental distribution, total and soluble copper distribution and preliminary

assessment of the leach response in small scale agitated tests. Subsequent stages can then

be planned on the basis of these results.

A series of column leach tests form the most important part of the testing programme. The two

key results to be ascertained are copper recovery and acid consumption. Operating parameters

such as leach cycle time, crush size, lift height, and acid addition methods can be optimised. A

program will often consist of a set of small column tests, to get a rough idea of the metallurgical

characteristics, followed by more detailed testing to optimise the conditions. A set of tall column

tests can then be conducted to confirm the results under conditions appropriate to a commercial

leach heap. Physical tests and water tests are also conducted to provide additional design

information.

For realistic predictions of full-scale production, column tests should closely model the

anticipated plant conditions. A set of laboratory procedures has been developed for this

purpose, and this paper includes a discussion of the technical aspects that must be considered

to ensure such resemblance. Suggestions for the presentation of results are also included,

along with a discussion of scaling up methods.

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3. TOTAL COPPER VS RECOVERABLE COPPER

Total copper assays do not normally indicate the material amenable to recovery by heap

leaching. The solubilities of the different oxide and sulphide minerals in acid/ferric solutions

vary greatly (Parkinson and Bhappu, 1995). Assay procedures have been developed so that

the copper grade can be expressed as `recoverable’ copper.

For carbonate/chrysocolla (`oxide’) leaching, a simple acid leach can give a reliable result. For

other oxide minerals, sulphide ores and mixed oxide/sulphide ores, which will be bio-leached, a

more complex method is required. The determination of the soluble components, and hence

the proportion recoverable by leaching, is called a diagnostic leach or sequential copper

analysis (Hiskey, 1997). This testing technique is based on oxide minerals being, in general

readily acid soluble, whereas bio-leachable (secondary sulphide) minerals are generally cyanide

soluble. In certain instances an acetic acid digest is also used to distinguish carbonate minerals

from chrysocolla and other oxides.

The basic procedure is to conduct a sequential cold acid, cyanide and aqua regia digestion to

establish the level of soluble copper. The sum of cold acid and cyanide soluble models the

long-term ferric/acid soluble species. The sequential method is important as species such as

cuprite and native copper are not totally CN soluble and need to be acid leach beforehand. The

CN test alone is also very difficult to get reproducible results from those species that have a

high acid solubility.

Leach recovery is best expressed as a percentage of soluble, or recoverable copper. In

defining the reserves it is also more meaningful to use recoverable than total copper. This gives

a better basis for economic analysis and engineering design.

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4. TEST PROGRAMME OVERVIEW

There is no one ideal test programme, every deposit is different and the programme will depend

upon the availability of sample and time. However, adopting a standard approach will promote

the most effective use of resources and generate the most reliable results. The programme

outlined is intended to give the best results in the shortest time, and make efficient use of the

samples. Factors affecting the test programme include:

• Size and grade of the ore body.

• Lithology (number of potential ore types).

• Mineralogy (Copper and gangue).

• Feasibility and cost of obtaining sample (May need to compromise between statistically accurate samples size and practically obtainable sample size).

• Project schedule.

• Financing requirements.

• Previous data available.

4.1 Preliminary Characterisation

After the ore types have been designated, a composite sample of each should be prepared.

Fine RC chips are adequate at this stage, although drill core may be preferable, and is

necessary in later stages. Multiple element analysis should be done to identify any components

potentially harmful to bio leaching or SX-EW, such as arsenic, manganese, molybdenum,

fluoride, silver and chloride.

Mineralogical examinations are very helpful for planning the column tests and interpreting the

results (Baum 1996). They should identify copper and gangue minerals, grain size,

intergrowths, associations etc. Later comparison of column leached residues against head

samples can support the leaching data, indicating the actual reactions occurring and the nature

of the copper in tails.

Small-scale sulphuric acid bottle roll tests are used to indicate the leach response of oxide

minerals. Similar tests using an acidified ferric solution can estimate the bio-leachable

components, including the secondary sulphides. Acid consumption and copper recovery are

determined by these tests, although acid use is generally much higher than in column tests.

Further bottle rolls, or shake flask tests using an inoculum obtained from site, or other projects,

can give a better idea of the bio-leachable component, and warn of any adverse effects the ore

may have on the bacteria.

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The leach conditions in a bottle roll test are very different from a commercial heap, so the

results have limited use. They can be considered as determinations of the maximum target

copper extraction and acid consumption. They are useful for giving quick estimates, and for

identifying the metallurgically significant rock types so that subsequent stages can be planned.

Cure tests can be performed to estimate how much acid should be added during agglomeration.

Acid consuming ores should be sufficiently cured so that the pH of the initial effluent is low

enough to continue leaching, but not so low as to waste acid. The tests involve curing ore

samples with various acid doses under conditions similar to an agglomerator. The range of

doses should be based on the maximum acid consumption determined by the bottle roll tests

(eg 25, 50, 75 & 100% of the maximum consumption).

4.2 Size Distribution

When drill core or bulk samples become available, it is possible to look at the crushing

characteristics and the resultant size distribution. The presence of clay or other fine material

may create problems for heap permeability and aeration. The distribution of soluble copper

across the range of particle sizes should also be understood.

4.3 Water Tests

A detailed site water analysis is an important aspect of the metallurgical assessment. The effect

of impurities on all stages of the process must be considered, particularly bio leaching and SX.

Where practical it is best to use site water for leach tests.

4.4 Column Tests - Programme Conditions

Heap leaching is a slow process, and column tests can take many months to complete. This

factor, along with scarcity of samples, tends to have a major effect on the planning of column

test programmes. The mineralogy is also a deciding factor, with different approaches taken for

sulphide and oxide minerals. Whether the minerals are finely disseminated or present in

fractures and veins, the presence of acid consuming gangue, the level of fine clay material, all

influence the test conditions selected. A typical programme consists of several stages, such as:

Phase I: Small columns (typically 2 m x 100 mm) using 12.5 mm or 25 mm crushed ore, to

estimate major leaching criteria.

Phase II: Small columns run under a variety of conditions, to optimise parameters such as

crush size and acid addition.

Phase III: Tall columns (typically 6 m x 150 mm) using the optimum leach method predicted

by phase II tests, to confirm leach response under realistic conditions.

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4.5 Optimisation

The parameters to be optimised tend to depend on whether the ore is oxide or sulphide

dominated. With oxide ore, the acid addition method can be critical to the acid consumption

(Miller et al, 1997). With sulphide ore the emphasis is normally on creating favourable

conditions for bacterial activity.

• Crush size: Single particle leach kinetics are improved by crushing finer, but physical

considerations such as permeability and aeration are favoured by a coarser crush. The

optimum size will be a compromise between these factors and the cost of crushing. Typical

sizes to test are p80s of 12.5 mm, 25 mm, and 50 mm. (Tertiary, secondary and primary).

Factors affecting the results include the mineral assemblage (veins/fractures or fine

dissemination), the importance of aeration, porosity of the host rock, and the copper grade.

• Acid addition method: In cases of high acid consumption, minimising the acid used can be

critical to the project’s economics. The pH should be kept low enough to continue copper

leaching, but not so low as to cause avoidable gangue consumption. A suite of tests using a

range of target feed acid levels can be conducted to see which results in the lowest overall

consumption without loss of copper recovery. The amount added in pre-treatment can also

be optimised. In the case of sulphide ores the leach/SX/EW process can be nett acid

generating, in which case external additions of H2SO4 should be minimised.

• Irrigation rate: This is usually 6 L/m2/hr, but can be increased to around 10 L/m2/hr for high

acid consuming oxides. Provided that sulphide ore is fully wetted, increasing the rate will

probably serve only to dilute the pregnant solution. In the event of aeration or short circuiting

problems, an intermittent application system may improve the leaching rate.

• Aeration: Simulating the air flow patterns in a leach heap is not easy to do in the laboratory.

To test the importance of aeration to a particular rock type there is an option of sealing the

column discharge from the atmosphere, or allowing air ingress at the base (Readett and

Miller 1997). Air sparging the feed solution helps ensure the bacterial activity in a bio-leach.

Forced aeration of sulphide ore is another option to consider.

• Grade effect on recovery: Throughout the life of a mine the grade of ore supplied to a heap

will probably vary, so it may be advisable to test the effect of different grades on leaching.

• Ore height: A trade-off exists between leach cycle time, which is reduced by lower heights,

and the cost of stacking and irrigation, which is higher for lower lifts. The choice is

particularly critical for sulphide ores, where poor aeration conditions at depth may control the

recovery. Column tests should ideally be done at the anticipated heap height, but shorter

columns can give adequate results if the chemical conditions, especially effluent pH, are

similar to the commercial operation (Dreier, 1995). A typical height is 6 m.

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4.6 Physical Tests

The following physical characteristics should be determined from column leach tests:

• Size distribution

• Slump = (Initial height - final height)/initial height x 100%.

• Agglomeration moisture requirement.

• Total moisture under active leach.

• Drained moisture.

• Drainage rate.

• Maximum percolation rate.

Additional testing should be undertaken to determine the following ore properties:

• Specific gravity.

• Bulk density.

• Moisture content.

• Angle of repose.

• Abrasion index.

• Compressive strength.

• Crushing work index.

If the ore contains considerable fine clayey material, permeability could be a major issue. In this

case additional percolation tests are called for.

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5. COLUMN TEST DESIGN

Column tests should be designed to simulate the conditions in a commercial leach heap as

closely as possible. This is not as easy as it sounds; particular attention should be paid to the

following:

• Water source.

• Initial solution (mature raffinate).

• Bacterial inoculum.

• Solution recycling.

• Temperature.

• Loading method.

• Acid and copper levels in column feed solutions.

• Aeration.

• Agglomeration.

• Sampling.

Open circuit tests have a limited application in preliminary tests on oxide ores. In all other cases

it is recommended to use a closed circuit with solvent extraction of copper from PLS and recycle

of raffinate. Closed circuit tests have advantages including:

• Concentration of ionic species builds up to equilibrium levels, giving more realistic estimates

of acid consumption and copper extraction.

• Bacterial populations can adapt to the conditions of the particular ore.

• Copper cations in the PLS are replaced by H+, giving a feed raffinate which replicates an SX

plant.

The disadvantages are that they are more time consuming, require extra equipment, and in the

early stages can produce highly acidic raffinates from the high grade PLS solutions.

5.1 Head and Tail Assay

To provide ongoing recovery estimates and give confidence in metallurgical accountability, an

accurate head assay should be determined for each column test. The recommended way to do

this is by size fraction (Keane, 1995). Head assays are done at each size fraction, along with a

composite assay to compare against the weighted average of the fraction assays. Ideally the

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head assay sample should be equal in weight to the column charge, but if the quantity of

sample is limited a reasonable compromise can be made.

After termination of the leach, the entire residue should be weighed, dried, screened, weighed

again, crushed and prepared for assay. As with the head sample, assaying by size fraction is

recommended for accurate results, and to enable copper recovery to be calculated at each size

fraction.

Wet screen analysis on the fine material is recommended for both head and tail samples.

This is a labour intensive procedure, but has numerous advantages:

• The size distribution of the sample can be examined to ensure it reasonably approximates a plant situation.

• A high level of metallurgical accountability improves confidence in the results.

• A reliable head assay enables recovery to be followed as the test proceeds, and interpretation of data can be done long before the final residue assays are available.

• Information regarding copper recovery by particle size is essential in optimising the crush size.

• The different sizing techniques (wet/dry) provide an insight into the ultra fine particle characteristics.

5.2 Inoculation and Agglomeration

Laboratory scale agglomeration can be conveniently done in a rotating drum such as a plastic

cement mixer. The agglomeration time should be similar to that of a commercial operation,

which is typically about 1 minute. Sufficient moisture should be added to dampen all of the ore,

but not leave a shiny surface. The amount of water required depends mainly on the fines

content, and should be determined by trial and error before hand. The acid cure, polymer and

inoculum should also be added during agglomeration if required.

Biological leach tests should preferable use solutions and bacteria from the specific site.

Alternative sources are similar operations, old workings and previous tests. An artificial raffinate

can be used, provided the full suite of chemical species and a sufficient bacterial population is

present. Leaching tests should always be conducted under the chemical and bacterial

conditions anticipated for the industrial scale operation, or the results will be misleading.

In the presence of swelling clays the physical benefits of agglomeration are particularly

important. Enough liquid must be added at this stage to satisfy the clays’ water of hydration and

prevent subsequent blinding of the ore.

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5.3 Solution Chemistry

In a commercial operation, freshly stacked ore will usually be a small proportion of the total ore

under leach. The rich solutions will be mixed with weaker solutions, thus stabilising both the

PLS copper grade and the raffinate acid concentration. This is not the case with column tests,

as a high PLS copper grade will produce a strongly acidic raffinate. There is therefore a need to

remove some acid from the solution inventory to avoid exaggerating the leach results.

In practice this means splitting the leach feed solution after the first SX cycle, and diluting it

down to a reasonable acid strength. The remaining strong solution can then be gradually added

back to the system. This has the disadvantage of diluting the other ions as well, unless dilution

is with the un-acidified initial solution. An alternative and preferred method is to partially

neutralise the solution to a target acid strength, using sodium hydroxide. The advantage of this

is that it avoids dilution of all the other ionic species, which may affect the leach. At all stages

the solutions must be carefully accounted for, to get an accurate acid consumption figure.

5.4 Solution Sampling

In closed circuit column tests, the sample volume should be kept to a minimum or it will interfere

with the leach accounting, and can complicate the acid consumption calculation. In the initial

stages daily sampling of PLS is recommended, but once conditions stabilise it is only necessary

to sample every few days. For convenience this can coincide with the batch SX cycle. Samples

(both feed and discharge) should be analysed for copper, free acid, iron (ferrous and ferric), pH

and Eh.

Late in the leach cycle a PLS sample should be subject to the same detailed characterisation

(ICP analysis) as the initial leach feed. A bacteria count may be advisable at this stage.

5.5 Temperature

In a commercial operation it is generally not possible to alter the temperature in the heap.

Laboratory tests should therefore be run as close as possible to the temperatures expected on

site. This temperature will be a function of the climatic conditions, and of the chemical reactions

occurring. Higher temperatures will exaggerate the leach kinetics, or may kill the bacteria in a

bio-leaching system.

5.6 Polymer Agglomeration

Ores with a high fines content create a challenge in terms of heap permeability. To deal with

this situation it is possible to add a polyacrylamide flocculant to the agglomerating solution. The

polymer should be compatible with the chemical conditions in the leach solutions, and should be

well mixed before addition.

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5.7 Aeration

Aeration within a full sized heap is particularly difficult to simulate in the laboratory, as oxygen

supply can often be the rate controlling factor. This is an inherent problem with column tests. A

leach heap is open to the atmosphere on all four sizes, and effectively sealed at the base. A

column test, however, is usually closed at the sides and open at the base. Depending on the

aims of a test, it may be desirable to either restrict or enhance the aeration. If excessive

aeration in a column could give exaggerated results, the base can be sealed from the

atmosphere by simply looping the discharge tube to create an air trap. To test the effect of

forced aeration, a low pressure air line can be connected near the base of the column.

In practice, leach feed solutions get well aerated while passing through the ponds, plant and

irrigation sprays. To replicate this in the laboratory, solutions should be left open to the

atmosphere, and possibly air sparged.

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6. PILOT LEACH HEAPS

The operation of pilot scale heap leach - SX - EW plants has been undertaken at several sites.

It is generally difficult to justify the expense of such a plant in terms of the metallurgical

information generated. The data can serve to confuse, rather than clarify the leaching

characteristics because of the inherent difficulties. The difficulties include getting an accurate

head sample, tolerance of metallurgical accounting and inflexibility regarding leach parameters.

A pilot plant can, however, be a very useful tool for staff training and transfer of knowledge in

areas unfamiliar with the technology.

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7. CALCULATION AND PRESENTATION OF RESULTS

7.1 Calculations.

A single column test can generate a very large volume of data. To facilitate interpretation of this

data, a standard reporting format is recommended. Following this approach will ensure that the

correct measurements are taken, and that the data can be readily understood and processed by

others. Particular attention should be paid to the following points:

• Final copper recovery and extraction curve based on calculated head, not assay head.

• Acid consumption as nett and gross.

• Daily measurement of volume fed and drained from the column.

In addition to leaching results, reports should contain head and tail size and copper distribution,

a summary of physical characteristics, and a log of batch SX and solution make-up. Producing

a summary sheet for each column is also recommended.

Principle calculations are done as follows:

• Copper recovery % = 100 x (Cumulative Cu in PLS - Cumulative Cu in Feed)/ Cu in

Head.

• Copper recovery % = 100 x (Cu in head - Cu in tail)/Cu in head.

Cu in PLS is the summation of the product of daily PLS volume (L) and assay (g/L). Cu in feed

is done the same way. Cu in head can be either assay head or calculated head. During the

course of a test the assay head is used to provide ongoing recovery estimates. At the end of

the test the calculated head is determined by adding the residual copper (tail) to the total

solution copper (PLS- feed).

The metallurgical accountability can be determined as calculated head / assay head. If the

head sample was properly split, and all assays were done accurately, the accountability should

be close to 100%.

There are two ways to calculate acid consumption. The easiest and most accurate way is to do

a mass balance over the closed cycle system:

• Acid consumption = Acid added to system - Acid removed from system - Acid inventory.

Acid added includes agglomeration and curing. Acid removed includes samples, and any

solution splitting and neutralisation. All these terms should include the acid equivalent of copper

(=1.54 x Cu grade). During the test, acid inventory can be estimated from the approximate

system volume and concentration. After the test, it is measured accurately once the final

drained solution has been extracted.

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The second method is a mass balance on the column only, and includes a credit for the acid

recovered by SX.

• Acid consumption = Acid in column feed - Acid in column discharge - 1.54 x Cu

extracted.

These two methods are for nett consumption and they should, in theory, give the same result.

The first method relies on weighed quantities of sulphuric acid, while the second uses the free

acid titration results. Results from the second method can be normalised at the end of a test to

conform with the first method. Acid consumption is usually expressed in kilograms per tonne of

ore, and can also be expressed as kilograms per kilogram of copper extracted.

7.2 Graphs

The following plots can enhance understanding of the leach response:

• Copper recovery over time (% of total and recoverable copper).

• Copper recovery over leachate flux (kL feed per tonne of ore) (% of total and recoverable copper) – Figure 1.

• Acid consumption over time (nett & gross) – Figure 2.

• Acid consumption over flux (nett & gross).

• Acid consumption against copper recovery (nett & gross) – Figure 3.

• ∆acid/∆Cu over time & flux.

• ∆acid/∆Cu against Cu recovery – Figure 4.

• Solution Cu grade over time (daily and cumulative).

• Solution Cu grade over flux (daily and cumulative) – Figure 1.

• Solution pH, Eh, Ferric/Ferrous over time, flux.

• Size distribution (Head and Tail).

• Copper distribution by particle size (head and tail).

• Copper recovery by particle size (% of total and recoverable copper).

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8. INTERPRETATION OF RESULTS

Scale up is an essential but often difficult part of the project development, and usually involves

empirical interpretation of the test results, and conservative safety margins. The graphical

techniques outlined above make the task considerably easier. Mathematical modelling

techniques have been developed and successfully applied to both oxide, sulphide and mixed

ores (Bartlett 1992).

The general shape of the recovery curve gives important indications about the nature of the

leaching. Following the changes in the solution chemistry is also essential. Comparison of the

size fraction head and tail assays shows how the leaching proceeds at different particle sizes.

With the exception of tenorite and cuprite, all oxide reactions occur very rapidly at ambient

temperatures. This is partly because oxidised copper minerals generally occur as fine grained,

paint-like coatings on fractures having very high surface areas. The driving force for these

reactions is the concentration of acid, so the rate is strongly pH dependent (Dreier, 1995).

Dissolution of the sulphide minerals is considerably slower than the oxides, with chalcopyrite

being particularly slow.

The presence of ferric is of critical importance to the indirect biological mechanisms. Iron is

released by the dissolution of pyrite, and is converted to its higher oxidation state, by the

bacterial biomass. The important point common to all sulphide reactions is the requirement of

oxygen.

Although acidic conditions are required, sulphide mineral dissolution rates are essentially

independent of pH. T. ferrooxidans can initiate growth at pH levels between about 1 and 5, with

the optimum for copper sulphide leaching being between 2.3 and 2.5. The ideal temperature

range is 28 - 35oC (Murr, 1980).

8.1 Leaching Models

For oxide dominated minerals, the dissolution step is considered rapid, so can be neglected in

modelling the rate. The rate is therefore controlled by the rate of diffusion of H+ ions into the

rock, and Cu2+ ions out of the rock. A sharp interface is assumed to form between the leached

rim and the unleached core, beyond which the lixiviant does not penetrate. Mathematically

modelling this system is relatively straightforward, based on the following relationship derived

from the quasi-steady-state reactant diffusion model:

1 - 2F/3 - (1 - F)2/3 = (2VCu Deff A0 / B r02 ) t (1)

Sulphide leaching kinetics are more difficult to model because the mineral dissolution rate is

significantly slow. The interface between the leached rim and unleached core cannot be

considered sharp, as the lixiviant (ferric) will penetrate into a partially leached section. Accurate

mathematical models have, however, been achieved, such as this mixed kinetics model:

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1 - 2F/3 - (1 - F)2/3 + (β’ / G ro)[1 - (1 - F)1/3] = γt / (G ro

2) (2)

here: β‘ = 2Deff / (Bβ) (3)

β= 6 ρore δ kz / (dm ρm) (4)

γ = 2MWm Deff c / (ρore B Ψ) (5)

Despite the complexity of this model, leaching data can be readily correlated by equation (2).

However, a unique solution cannot be found unless either the diffusion rate or the reaction rate

is known. If not, the model contains too many degrees of freedom to be correlated against

leach data alone. A reasonable estimate must therefore be made of effective porosity and the

diffusion coefficient, or mineral surface area and the chemical rate constant. The relative effects

of diffusion and reaction kinetics can then be compared over the course of the leaching cycle.

The leaching rate is always chemically controlled in the early stages, before the leached rim

grows into a significant diffusion barrier. Typically though, diffusion is the rate-controlling step

over the majority of the leaching cycle.

8.2 Macroscopic Leaching Kinetics

The particle leaching models described above can be modified to apply to large commercial

leach heaps. The approach taken again depends on whether the ore is considered sulphide or

oxide dominant. This is because when bio-leaching sulphide ores the leaching agent, ferric, is

able to be regenerated throughout the ore mass continually, provided there is enough oxygen to

maintain bacterial activity. In contrast, in acid leaching of oxide ores the reagent is consumed

as it passes through the ore, resulting in a vertical concentration gradient.

The model for acid leaching of oxide ores described by equation (1) has a serious limitation:

The feed acid concentration, Ao, is considered constant. This approximation may be valid for

thin layer leaching, but not for a significant heap height. For the model to apply, the heap must

be considered as a series of horizontal layers, with copper concentration increasing and acid

concentration decreasing down through the heap. The recovery is calculated at each level,

during each time period, and for each particle size by solving the diffusion equation and mass

balance for each layer (Bartlett, 1992).

Until the acid front has propagated through the heap, no copper will be leached from the lower

levels. In these early stages the recovery rate is closely related to the acid addition rate, so

diffusion effects will be largely irrelevant. If the ore has been cured, the availability of acid at the

mineral surface is less likely to be rate controlling, but there will still be a significant vertical acid

gradient. This is especially the case for high acid consuming gangue material, carbonate ores,

and in the early stages for chrysocolla ores.

The model for ferric leaching of sulphide ores described by equations (2), (3), (4) and (5) is

based on the assumption that there is an adequate supply of ferric throughout the heap. In

practice this may not be the case, as there are two other potentially rate-limiting factors (Murr,

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1980). One is the supply of air, the other is the bacterial conversion of ferrous to ferric. Natural

air convection is driven by temperature gradients throughout the heap, which are created by the

exothermic dissolution reactions. If the bulk voidage is insufficient, air permeability will be

restricted and parts of the heap will be starved of oxygen and carbon dioxide. Excessive fines

can be the cause of this. The rate of iron oxidation can also be restricted by insufficient total

iron, the presence of toxins, and temperatures outside the operating range of T. ferrooxidans. A

ferric limited regime should be suspected when the recovery curve is linear with respect to time,

rather than following the more usual curve.

The models and methods presented above treat sulphide and oxide ores separately, but in

practice this may not be possible. Transition zones are frequently encountered containing a

variety of minerals, in which case the interpretation techniques must be tailored to suit each

unique situation.

8.3 Acid Balance

The overall acid balance of an ore depends on a complex set of reactions, which can be broadly

classified as acid consuming, producing, or buffering (Templeton and Schlitt, 1997). Irreversible

reactions either consume or generate acid, while reversible reactions have a buffering effect.

The reactions include copper dissolution, iron and sulphur oxidation, gangue interactions,

precipitation and solvent extraction. Strategies to minimise acid consumption over the life of

mine can have significant economic consequences, particularly for oxidised ores.

Acid additions in both the pre-treatment and leaching stages are varied in order to identify the

optimum in terms of acid consumption, leach kinetics and overall recovery. Blending of different

ore types may also have benefits. The fundamental aim is to balance the conflicting aims of

maximum copper extraction and minimum gangue consumption.

In general, acid consumption increases with finer crushing, lixiviant-ore contact time, and feed

acid concentration. It follows that column tests should be done at the same size distribution,

irrigation rate and height as the commercial heap, if the acid consumption is to be applied to

scale-up. Tests should be done in closed cycle to allow soluble impurities to build up to

equilibrium levels, creating a buffering effect.

To decide upon the optimum acid concentration and application method, the results from

several tests need to be compared. Plots of cumulative acid consumption versus copper

recovery can help identify the optimum, Figure 3. They can also indicate the point, late in the

leach, where the cost of additional acid exceeds the value of the additional copper extracted,

and further irrigation becomes uneconomic, Figure 4. This method can also indicate if there is

any difference between the ‘ultimate’ acid consumption, at the end of economic leaching,

between different acid addition systems.

19

9. SCALE UP

The rigorous mathematical models described previously are rarely used in practice. The

fundamental deterministic data are generally not available, and if they are it is generally difficult

to justify the effort required in setting up the tests to get the data. Alternative empirical models

provide methods that are easy to use, and can be applied in most situations. These take

advantage of the way that leach curves tend to conveniently follow simply mathematical

relationships such as:

F = a + b ln Q (6)

While leach data may closely fit this form of this equation – Figure 1, extrapolation is risky

because the equation is not bounded, and can predict recoveries greater than 100%. This can

be avoided by using a transformed style:

ln (1-F) = -K Q (7)

This equation is bounded at F = 1, but generally will predict higher recovery rates than equation

(6) below about 98% recovery. A modification of equation 1, combining several parameters into

a single constant, can also be used to correlate test data (Bartlett, 1997):

1 - 2F/3 - (1 - F)2/3 = K Q (8)

An example is shown in Figure 5.

In these equations, time can be used as a parameter instead of lixiviant volume, provided the

irrigation rate is constant. When applying short column results to a higher heap, it is important

to consider recovery as a function of volume of solution per tonne of ore rather than time.

The complexities and uncertainties inherent in a commercial heap create a challenge in

interpreting test results. All of the available models are based on certain idealised assumptions,

the validity of which should be reviewed before putting too much faith in them. Laboratory

results need to be treated with caution because full-scale heaps are rarely placed with the same

care as in columns. Uneven loading may create solution channelling, surface ponding, internal

ponding and fines mobilisation, and dry patches within the heap. As a result it is usual to

discount the recovery to take some account of the non ideal placement in the field heap.

One method that has been adopted is to discount the recovery by –5% at a selected leachate

flux (usually corresponding to 85% recovery of the soluble copper). This modest reduction

when applied to the log model has the effect of increasing the leach time by 45% to achieve the

original recovery level. As such the technique is showing promise in the ability to provide a

measure of comfort in the predicted field recovery estimates.

The acid consumption associated with these increased field leach times is a more difficult

parameter to predict. It is important that acid consumption be tested on the actual field selected

20

heap height. Some sample data from a couple of tests shows the significant effect that the

height has on the measured consumption rate.

Column Height 2m 6m

Acid consumption rate kg/t/day @ 6g/L acid 0.453 0.184

@ 3 g/L acid 0.144 0.089

A number of workers have emphasised the importance of obtaining data on acid consumption

form tests that model the field situation.

“..the experimental determination of gangue acid consumption should replicate the commercial

crush size distribution of the ore and the heap irrigation rate. Column or test heap height should

also match that of the actual heaps, and the leach solutions should come from long-term,

locked-cycle tests in which impurities have built up to steady state values.” (Templeton and

Schlitt, 1997)

It has been found that even with acid agglomerated materials, the acid consumption (once all

the copper equivalence has been sorted out) is linear with time, since consumption continues

during those periods even when irrigation is halted, Figure 2.

Consumption rate is also related to the concentration of application, with higher strengths

having high consumption rates and vica versa. It is thus important to consider the leaching

management plan when modelling the acid consumption as it is likely that a two stage leach will

be conducted with low acid concentration solutions late in the leach cycle. This allows the

modelling to account for the lower consumption rate in the last half of the recovery period.

In many cases the acid consumption will be limited by the application rate of acid to the heap

and the tested consumption rate will exceed the physical supply of acid when extrapolated to

deeper lifts. In these cases where the model acid consumption is in excess of the supply needs

the total consumption must be limited to the acid availability. In the latter stages of leaching

where the ∆acid/∆Cu ratio is > 3:1 there is negligible effect on the copper dissolution kinetics

with a reduction in the acid concentration.

21

10 CONCLUSIONS

The testing of ores for heap leach amenability is a relatively complex one that needs proper

planning of the programme and outcomes.

With properly executed tests the information can be obtained to properly design and specify the

heap leach and solvent extraction sections of the process plant.

Modelling of the leaching extraction curve has been accomplished recently with three different

models that are of increasing complexity. Extrapolation form the test data needs to be done

with one ore more of these models to ensure that the result does not exceed the leachability of

the ore.

Modelling of the acid consumption is a more complex problem as the acid concentration profile

changes at the leachate moves down the column. At this stage testing at the proposed lift

height gives the most reliable results for use in the prediction of the overall acid consumption.

There is scope for modelling the acid consumption using different acid concentrations in

appropriate parts of the leach cycle. This allows high consuming ores to be leached with the

minimum acid use. Modelling of this requires knowledge of the different acid consumption rates

with differing acid concentrations.

22

11. ACKNOWLEDGEMENTS

The views expressed in this paper reflect those of the authors and do not necessarily reflect

those of other interested parties. The work is a culmination of many small progressive steps in

the understanding of the leach process and the factors that control it. The experience gained on

a number of projects is acknowledged gratefully to the Egis clients that have contributed their

faith in us.

23

12. NOMENCLATURE

A0 feed acid concentration

a and b constants

B reaction stoichiometry coefficient (moles acid per mole copper)

c oxidant concentration (eg Fe3+)

Deff effective diffusivity

dm average mineral grain diameter

F fractional copper recovery

G ore grade (mass fraction of mineral)

K constant

kz reaction rate constant

MWm mineral molecular weight

Q volume of lixiviant applied per unit mass of ore

ro particle radius

t leach time

VCu rock volume containing one mole of copper

δ thickness of the reaction zone considered during ∆t

ρm mineral gain density

ρore ore density

ψ sphericity factor (= surface area of a same diameter sphere / actual surface area)

24

13. REFERENCES

BARTLETT, R. W. Solution Mining, Leaching and Fluid Recovery of Materials, Gordon and

Breach, New York, NY, 1992

BARTLETT, R. W. Metal Extraction from Ores by Heap Leaching, Metall. Trans., vol. 28(B),

1997. pp 529 – 545.

BAUM, W. Optimizing Copper Leaching/SX-EW Operations with Mineralogical Data, SME

Annual Meeting, Phoenix, 1996.

DREIER, J. Geochemical Aspects of Copper Heap Leaching, SME Copper Heap Leach Short

Course, Denver, 1995.

HISKEY, J. B. Diagnostic Leaching of Copper Bearing Minerals. Heap and Dump Leaching,

AMF Course 119/97, Newcastle, 1997.

KEANE, J. M. Commercial Ore Testing, SME Copper Heap Leach Short Course, Denver, 1995.

MILLER, G. M. and DICKINOWSKI, W. and STUART, M. Heap Leach Testing and Operation of

the Mt Cuthbert Sulphide Ores, International Biohydrometallurgy Symposium IBS97 - Biomine

97, AMF, Sydney 1997.

MURR, L. E. Theory and Practice of Copper Sulphide Leaching in Dumps and In-Situ, Minerals

Sci. Engng., vol. 12, no. 3. July, 1980. pp 121 – 189.

PARKISON, G. A. and BHAPPU, R. B. The Sequential Copper Analysis Method – Geological,

Mineralogical, and Metallurgical Implications, SME Annual Meeting, Denver, 1995.

READETT, D. and MILLER, G. M. Engineering and Process Developments Associated with

Industrial Scale Copper Bio-leaching, International Biohydrometallurgy Symposium IBS97 -

Biomine 97, AMF, Sydney 1997.

TEMPLETON, J. H. and SCHLITT, W. J. Method for Determining the Sulfuric Acid Balance in

Copper-Leach Systems, Minerals and Metallurgical Processing, August 1997.

25

FIGURES

Figure 1. Copper Recovery and PLS Grade vs Flux inc recovery log model.

Figure 2. Acid Consumption vs time inc linear model of consumption rate.

y = 0.1382Ln(x) + 0.6269

R2 = 0.9947

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

0 1 2 3 4 5 6 7 8

Leach Flux kL/t

Recovery F (%)

F

Log. (F)

y = 0.1146x + 4.5692

R2 = 0.9978

y = 0.2103x - 0.7011

R2 = 0.9944

y = 0.1703x + 18.403

R2 = 0.9977

y = 0.1584x + 19.279

R2 = 0.9945

0

5

10

15

20

25

30

35

40

45

40 60 80 100 120 140

Leach Days

Net Acid Consumption kg/t (solution calcs)

C1 KNS 15mm

C2 KNS 10mm

C3 KN 15mm

C4 KN 10mm

26

Figure 3. Acid Consumption vs Copper Recovery.

Figure 4. ∆acid/∆Cu ratio vs Copper Recovery.

0

5

10

15

20

25

30

35

40

45

0 20 40 60 80 100

Copper Recovery %

∆ Acid/ ∆ Copper Ratio

0

5

10

15

20

25

30

35

40

0 10 20 30 40 50 60 70 80

Copper Recovery %

NettAcidConsumptionkg/t

C1 KNS 15mm

C2 KNS 10mm

C3 KN 15mm

C4 KN 10mm

27

Figure 5. Lumped Parameter Diffusion Model of Extraction

•

y = 0.0332x + 0.0288

R2 = 0.9818

0%

2%

4%

6%

8%

10%

12%

14%

16%

0.0 0.5 1.0 1.5 2.0 2.5 3.0 3.5 4.0

Leach Flux kL/t

Calculated reduced Recovery

1-2/3(R)-(1-R)^(2/3)

Linear (1-2/3(R)-(1-R)^(2/3))