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Reducing Selectivity in Narrow-Vein Mining through the Integration of Underground

Pre-concentration

Andrew Bamber, Bern Klein, Mario Morin, Malcolm Scoble

 Department of Mining Engineering, University of British Columbia, Vancouver, Canada

Abstract

The introduction of pre-concentration technology into the underground narrow-vein mining scenario has the

 potential to change the entire economics of utilizing bulk-mining methods in these situations. An idealized

integrated underground mining and processing system, comprising elements of bulk mining, underground pre-

concentration and backfill systems is presented. Through simulation and a systems engineering approach, theimplications of enabling the utilization of bulk mining methods in a narrow vein deposit through the adoption of

this technology is considered. Parameters considered include orebody characteristics, ore mineralogy, grade,

metal value, mining rate and cost.

Keywords – narrow-vein mining, bulk methods, underground pre-concentration, backfill, economics

1.  Introduction

The majority of Canadian hard-rock narrow-vein mines are mature, and face the need to mine deeper deposits inthe face of increasing costs and declining grades. Obstacles to the successful mining of such deposits include

ground control, material handling and ventilation. In addition to these issues, the mining of narrow-vein deposits

faces additional economic challenges in the low productivity of the typical narrow-vein mining method, the

management of dilution and therefore grade, and the optimal integration of fill into the mining cycle.

UBC has been involved with INCO in a strategic research initiative into mine-mill integration, and more

specifically underground pre-concentration, since 2000. Several enabling technologies such as coarse-particle

mineral processes, hydraulic transport and modular cemented backfill systems have been identified, andintegrated using systems engineering techniques into a conceptual underground mining and processing system

for the rejection and disposal of waste. The grade of ore delivered to surface is substantially increased through

the rejection of barren waste underground. Previous research has indicated that substantial operating and capital

cost savings can be achieved, as well as a reduction in the surface footprint of the operation.

A case study has been undertaken for Falconbridge’s Onaping Depth Deposit in Sudbury, Ontario to assess the

 potential impacts of underground pre-concentration. Geological research and mineralogical evaluation of typicalSudbury Igneous Complex ores indicate that between 20 – 70% of the ore mined can be rejected with a good

metallurgical recovery. Several coarse-particle separation technologies have been identified and tested for use in

the underground process plant. Waste thus rejected is highly competent, and is suitable for use as a source of

aggregate for cemented backfill of superior mechanical properties. A conceptual process plant has beendeveloped and integrated into the underground mining and material handling system. Several impacts of the

implementation of underground pre-concentration on the operation have been identified, including:

•  Lowering of the cost of metal production through rejection of up to 60% dilution underground atgreater than 95% metal recovery

•  Savings in underground haulage, hoisting, surface transport, milling and tailings disposal

•  Lowering the cutoff grade through the reduction in operating costs

•  Increasing the mineral reserve though lowering of the effective cutoff grade

•  Decrease in selectivity of mining, thus increasing productivity•  A possible increase in minimum stoping height (ground conditions permitting), facilitating the use of

more productive mining equipment and the introduction of automated roof-bolters

•  Potential change in mining method to bulk-mining techniques due to the inclusion of a continuous,efficient and effective waste-rejection step within the mining cycle

•  Improvement in ground control and a potential reduction in rockbursting due to the introduction ofsuperior backfill into the mining void.

A discussion of the integration of the mining, processing, and backfill systems into the case study mine is

 presented. Revisions to the present mining method which will increase the productivity and lower the costs of

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the method are proposed. Conclusions to the feasibility and benefits of implementing such an integrated mining

and processing system will be presented.

2.  Impact of introducing Bulk Mining Methods in Narrow Vein Situations

Consider a deep, steeply-dipping, narrow tabular orebody as shown in Figure 1. The orebody is deep, and thus a

low cost of mining is critical to the feasibility of exploiting this resource. Due to the geometry of the orebody,

typically a narrow-vein mining method such as longwall stoping or drift-and-fill would be preferred.

Figure 1 – Example of Deep, Moderately Dipping Narrow-vein Orebody Configuration

Let it be assumed that a drift-and fill method has been selected, orebody access is via a central footwall drift,and two strike-oriented drift-headings are driven simultaneously from the central drift out towards the East and

West abutments. A typical narrow-vein drift and fill panel on strike can be represented as shown in Fig 2. Seamthickness ‘t’ is 1m. In the interests of minimising dilution, panel dimensions outside of the seam must be

minimised, thus panel dimensions for a seam width of 1m would be of the order of 3m x 2.5m, with a round

length of approximately 2 – 2.5m. The seam angle ‘a’ in this case is 55 ˚. This panel dimension would besuitable for a mining suite comprising a single-boom jumbo, a 4.5m3 LHD and a roof-bolter.

Figure 2 – Typical Narrow-vein Drift-and-fill Panel on Strike

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The overall characteristics, including the grade of the round can be determined from the dimensions andgeometry of the panel, as well as the density and grade of the ore and waste respectively. The critical parameters

in this calculation are presented in Table 1.

Table 1 – Orebody Characteristics for Grade Determination

Variable Description Value

Ore thickness t 1mOrebody dip a 55deg

Ore grade g 10%

Ore density ρo  3.4

Waste grade w 1%

Waste density ρw  2.8

The grade of the round overall can be calculated based on the seam thickness and orientation, panel dimensions,

and the relative grade of the ore and the waste.

The quantity of metal in the round can be calculated as follows:

Ore mass mo = ρotwd (1)cos(a)

Ore grade = g

Waste mass mw  = ρw(wdh – twd ) (2)

cos(a)

Waste grade = w

Thus overall panel grade can be calculated according to the ratio of the mass of ore and waste in the round:

 p = mog + mww (3)mo + mw

For the parameter values in Table 1, the grade of a narrow seam drift-and-fill round on this orebody would be

From (1) mo  = 35.57 t and from (2) mw = 12.71 t

Dilution in the round is thus 35%. Substituting for mo, g, mw and w in (3), the overall grade of the round is

 p = [35.57(10) + 1(12.71)]/(48.28)= 7.63%

Productivity for this type of equipment suite is low, and therefore unit mining costs for narrow-seam methods

are high. The unit cost of mining can be lowered by increasing the productivity of the mining suite. This will

entail an increase in the panel dimensions, and consequently a decrease in the overall panel grade, and thus theoverall as-mined grade.

Panel dimensions for a mining suite comprising a double-boom jumbo, 6m3 LHD and an automated roof-bolter

would be of the order of 6m x 3m, with a face advance of 3m. The grade of the round as calculated by the abovemethod would be:

mo  = 106.7 t

mw  = 63.3 t

Dilution of the round is now 59.3%. The grade of the ore and waste is unchanged, thus the overall panel grade p

would be:

 p = 6.65%

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The grade of ore generated from mining operations is decreased by 13% through a 41% increase in panel size;

this figure would be greater for a greater increase in panel dimensions. However, mining costs arising from thismethod will also be lower due to the higher productivity of the suite. The value of the ore delivered to surface is

dependent on the grade of the ore, as well as the price of the metal; if this decrease in value is greater than the

cost saving of adopting the more productive method, then this approach is unfeasible.

In the interests of clarity and simplicity, a tabular narrow vein deposit has been chosen. However, Figure 3

shows a number of other presentations of ore or “panel mineralogies” which could be considered narrow vein,and for which the volumetric and therefore grade calculations are substantially more complex.

Figure 3 – Potential Panel Mineralogies Found in Narrow-vein Situations

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While the panel mineralogies in each case differ, due to the narrow-vein nature of the deposit, increasing the

mining width consequently decreases the grade of ore mined. However, in each case, there is a substantial proportion of the tons mined which can be distinguished from the ore and rejected as barren waste.

3. Impact of Underground Pre-concentration on Ore Tonnage, Grade and Mining Costs

UBC is involved in a strategic research programme into underground pre-concentration. It has been identifiedthat savings in haulage, hoisting and surface milling can be achieved through the rejection of waste

underground. The grade of ore delivered to surface can be improved by up to 100% through the rejection of

gangue material underground in an integrated underground mining and processing system (Bamber et al, 2004).

A case study has been undertaken in order to identify suitable design criteria for a typical underground pre-

concentration facility. Basic design criteria identified through research and testwork are:

•  Waste must be rejected as coarse and as early in the mining cycle as possible

•  Process plant must be compact, robust, and flexible in the face of variations in feed size, tonnageand grade

•  Coarse-particle processing technologies are preferred

•  Maximum recovery of the economic mineral(s) must be achieved

•  A maximum of 60% only of the tons mined can be rejected

•  Waste products must be suitable for use as backfill

Coarse particle technologies that have been identified and tested thus far include:

•  Optical sorting

•  Conductivity sorting

•  High intensity magnetic separation

•  Gravity concentration

•  Dense media separation

Coarse-particle grinding and flotation has also been identified as a viable technology, but has not as yet been

tested. A basic process flowsheet of the integrated mining and processing system is shown in Figure 4.

Figure 4 – Integrated Underground Mining and Processing System

The optimum performance of an underground pre-concentration facility has been determined to be 60% waste

rejection at 95% metal recovery. The overall metallurgical balance for such an underground pre-concentration plant is shown in Figure 5.

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Pre-concentrator 

capacity t

mass pull 40%

recovery 95%

Feed

tons F

grade g

dilution d

Concentrate

tons C = F - T

grade =

0.95*(Fg)/(F-T)

Rejects

tons T = 0.6F

grade =

0.05*(Fg)/T

 Figure 5 – Metallurgical Balance for an Underground Mining and Processing System

Feed is processed at as coarse a size as possible, typically 100% - 80mm. Waste rejects are slightly coarser than

 plant feed due to the characteristic increase in grade towards the finer fractions exhibited by massive-sulphide base metal ores. Concentrate grade is approximately double the feed grade due to the rejection of barren waste.

Metal values in the rejects are low due to the high metallurgical recovery of the processes that have been

researched.

There are a number of impacts of implementing such a pre-concentration system on the underground mining

system. In the example cited in section 2, the panel grade resulting from the adoption of bulk mining techniqueswas shown to be 6.65% with 59% dilution. Assuming a mining rate of 3000 tpd, pre-concentrated ore at 1800

tpd and a grade of 10.53% could be delivered to surface through the adoption of underground pre-concentration.

Waste products are suitable for use as fill either as rockfill, or with further processing as high-strength cemented backfill. Implementation of the system was shown through modelling to reduce mining and milling costs by

20%. A further benefit that would be enjoyed in a greenfields situation would be a reduction of 10% in the

initial capital outlay due to a reduction in tons, and therefore the size of the hauling, hoisting, milling andtailings disposal facilities at the mine.

4.  Case Study – Onaping Depth

For the purposes of this paper, the authors have chosen to examine the impact of pre-concentration on the

extraction of Falconbridge’s Onaping Depth deposit, located in the Sudbury Basin in Ontario. In the late 1990’s,

Falconbridge was considering the mining of this resource but has chosen to postpone its developmentindefinitely. The safe extraction of the ore at these depths has proven itself to be too risky at this time due to

ground stability issues.

4.1.  Geology, Mineralogy and MiningOre deposits of the North Range of the Sudbury Igneous Complex can be classified into 5 generic ore types

(Coats and Snadjr, 1984):•  Disseminated sulphides in the mafic norite of the Complex

•  Disseminated sulphides, stringer sulphides, massive sulphide lenses and interstitial sulphides in breccias in the sublayer norites

•  Massive sulphides and interstitial sulphides in breccia at the contact of the sublayer norite and theFootwall rocks

•  Sulphide lenses associated with massive sulphide stringers in Footwall rocks

•  Cu-rich massive sulphide stringers and dykes in Footwall rocks

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Ores of the Craig / Onaping occur as two of the above ore types. The Main or ‘contact’ ore zones are shallower

deposits comprising massive, disseminated and breccia sulphides occupying a deep embayment in themineralized zone of the Footwall breccia of the SIC. The Main ore zone sulphides grade into Footwall zone ores

comprising Cu-rich massive sulphide stringers varying in width between 0 – 6m. It is thought that these ores

originate from copper remobilization in the contact ore zone. Sulphide stringers in the SIC are typically oriented parallel to the Footwall contact, and are thought to occupy stress-related fractures in the footwall normal to the

compressive stress generated by the igneous intrusion. The brecciation at the contact between the sublayer

norites and the footwall rocks is related to shock generated by the rate of the igneous intrusion. Principal North-Range sulphides are chalcopyrite, pentlandite, pyrrhotite, and pyrite. PGM’s occur as sperrylite or as a solid

solution in the chalcopyrite or pentlandite.

Principal characteristics of the Onaping Depth orebody are shown in Table 2 (Swan & Campbell, 2001).

Table 2 – Orebody Characteristics of Onaping Depth Deposit

Property Value

Ore type Massive and stringer sulphides (pentlandite / chalcopyrite) in

Sudbury Breccia

UCS 20 GPa

Hangingwall Sudbury Breccia

UCS 100 GPa

Footwall Footwall gneiss

Dilution 30%Extraction 80%

Depth 2300 – 2600 m

Thickness Variable 60 – 100m, average 70m

Dip 30˚ 

Strike 420m

Total Resource (Indicated + Inferred) 15.8 Mt

Cu 1.16 %

 Ni 2.6 %

TPM (estimate) 0.65 g/t

 Narrow vein orebodies are generally classified as seams of between 0.1 and 3m with an intermediate or steep

dip of 20 - 55o

(Miller-Tate et al, 1995). The Onaping Depth orebody, like the Strathcona Copper Zone and the

McCreedy East footwall ores, comprises multiple stringers of varying thickness within the thickness of theorezone, but can be considered a narrow-vein orebody due to the narrowness and steep inclination of the

individual veins. The mining method in use at Strathcona and McCreedy on these ores is drift-and-fill, which is

a suitable narrow-vein method.

Based on the geometry, depth and orientation of the orebody a number of mining methods are suggested (De

Souza et al 1987):

•  Longwall stoping

•  Drift-and-fill

•  Cut-and-fill (overhand or underhand)

•  Shrinkage stoping (wider veins with steep dip only)

•  Sublevel caving (wider veins and steep dips in competent hangingwall only)

In each case, resueing may also be adopted in addition to the basic method in order to reduce dilution. A second

factor to be considered in selecting the mining method is the geotechnical characteristics of the ore, footwall and

hangingwall rocks. This can be defined by the Q index (Barton, 1988) as follows:

Q = (RQD) Jr Jw (4)

Jn Ja (SRF)

Where: RQD - Rock Quality Designation

SRF - Stress Reduction FactorJr - Joint roughness number

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  Jw - Joint water reduction number

Jn - Joint set numberJa - Joint set alteration number

Other mining method selections based on orebody geometry and depth as well as the geotechnical setting of thedeposit are similar and include the Nicholas method (Nicholas, 1981) and the UBC modified Nicholas approach

(Pakalnis et al, 1995). A low Q index indicates that low-volume supported methods such as longwall stoping or

drift-and-fill methods would be preferred. A moderate Q index indicates a preference for higher volumemethods such as room-and-pillar, cut and fill or shrinkage stoping. A high Q indicates enables the consideration

of high volume methods such as open stoping, VCR or sublevel caving.

Applying all method selection criteria, drift- or cut-and-fill methods are preferred in conditions of high stress, or

 poor ore, hangingwall and footwall strength or condition through the improved ground control offered through

the use of fill. The benefits of increased ground control in this method can be further improved through the useof higher strength fills, although this is not generally considered economic.

In the case of Onaping Depth, underhand cut-and-fill was the method selected due to concerns for rockbursting precipitated by the high level of rock stress at this depth, and the brittle nature of the Sudbury Breccia (Romer,

2001). Principal access to the orebody was to be via a sub-shaft developed from the 4500 level of Craig Mine

(see Figure 6).

Figure 6 – Onaping Depth – Location and Access (www.falconbridge.com/projects)

A number of options for secondary access to the orebody were considered. These included:

•  Strike oriented mining with hangingwall access

•  Strike oriented mining with footwall access

•  Strike oriented mining with east abutment access

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•  Dip-oriented mining with hangingwall access

•  Dip-oriented mining with footwall access

•  Dip-oriented mining with east abutment access

The final mining method selection was dip-oriented sequential underhand cut-and-fill panel in a weave pattern

with orebody access obtained from a strike-oriented footwall drift (Figure 7). A sequence of two underhand cut-

and fill levels with alternating sill-cut levels was used in order to further minimize anticipated the rock-bursting problem. Based on a reserve estimate of 16Mt, a mining rate of 2933 tpd over 341 working days / year gives a

life-of-mine of 16 years.

Figure 7 – Typical Onaping Depth Drift-and-Fill Mining Level (Romer and Lawler, 2001)

4.2. Evaluation of Cut-and-fill Mining with Underground Pre-concentration

It has been demonstrated that underground pre-concentration can decrease the tons, and increase the grade of ore

delivered to surface (Bamber, 2004). An estimate of capital and operating cost savings is given for the case

study outlined above, based on the mining of the Onaping Depth orebody alone, and delivering the ore to adedicated mill nearby.

Impacts on the tonnage and grade of ore delivered to surface, based on a mining rate of 2933 tpd and themetallurgical performance indicated above is shown Table 3.

Table 3 – Comparison on Impacts Caused by Pre-concentration on Onaping Depth Ore

In situ After Pre-concentration

Mining Rate (tpd) 2933 1760

Dilution (%) 30 0

Cu grade (%) 1.13 1.8

Ni grade (%) 2.6 4.12

PGM’s (g/t) 0.64 1.01

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4.3 Capital and Operating Cost Impacts

Capital cost impacts are estimated based on the reduction of tonnage and therefore size of downstream facilities

after the rejection of waste. Capital impacts are calculated using the industry capital estimation formula that can

 be used to estimate the cost of an envisaged facility from the cost of an existing similar facility, according to therelation:

C2 = C1 (t2/t1)

2/3

  (5)

Where

C2 = new capital costC1  = existing capital cost

t1  = throughput of existing facility

t2 = throughput of new facility

In this way the theoretical capital saving enabled through the rejection of the waste fraction can be calculated.The mining rate is unchanged; therefore there is no change in the capital requirement for mining. Capital

savings are enjoyed in haulage, hoisting, surface transport, and milling and tailings disposal operations. The

additional capital cost of the new pre-concentration facility must be accounted for in this estimate: if the cost ofthe pre-concentration facility is greater than the capital saving, there will be an increase in the overall capital

required for the system. Capital savings for scenarios with low proportions of waste rejection are ofteninsufficient to offset the increase in capital for the pre-concentration facility. Based on industry averages for

Capex in mining, underground haulage, hoisting, and surface milling, the break-even point is approximately

30%, thus it can be seen in the example of Onaping Depth that there is a slight increase in capital cost, based onestimated figures for the capital cost of the baseline system. Substituting into equation 5 for the estimated

Capex, we obtain the results shown in Table 4.

Table 4 – Capex Cost Comparison for Medium-Depth Orebody without/with Pre-concentration

Mining Hauling

Pre-

concentration Hoisting

Surface

Transport Milling Tailings Total

15 10 0 25 10 35 5.00 100

15 8.39 20 20.98 8.39 29.39 4.19 106.36

Operating cost savings can be calculated in two ways. In the scenario described in this paper, the possibility of

facilitating cut-and-fill instead of drift-and-fill methods through the implementation of underground pre-concentration results in a reduction of mining costs from those typical for drift and fill to those more typical of

cut and fill. This can be quantified through historical data for similar mining operations from sources such as the

Canadian Mines Handbook. A comparison of industry average costs for drift-and-fill methods results in thefollowing estimate of typical mining and milling costs for a medium-depth drift-and-fill operation (see Table 5).

Table 5 – Mining and Milling Costs for Medium-Depth Drift-and-Fill Operation

Mining

Mucking/

haulage Backfill

Crush/

Convey Hoisting

Surface

Haul Milling Tailings Total

46.90 11.70 24.28 1.66 1.90 4.45 7.34 2.00 100.25

Comparing costs for drift-and-fill to cut-and-fill from the same source indicates the type of saving possiblethrough the facilitation of the higher productivity, lower cost method (see Table 6). It must be pointed out that

for Onaping Depth, underhand cut-and-fill was chosen as the only practicable and safe mining method so the

cost comparison and related conclusions may not be entirely valid.

Table 6 – Mining and Milling Costs for Medium-Depth Cut-and-Fill Operation

Mining

Mucking/

haulage Backfill

Crush/

convey Hoisting

Surface

Haul Milling Tailings Total

25.15 7.74 8.24 3.33 1.53 4.45 7.34 2.00 59.78

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A possible operating cost saving of 40% is indicated in this instance. Operating cost savings can also be enjoyed

within the same mining method through the reduction in tons processed in the downstream activities. The savingdue to the reduction in tons applies only to the variable portion of the operating costs, fixed operating costs

remain unaffected by this reduction. It is assumed that unit-operating costs do not vary significantly over small

variations in the feed tonnage to a particular process. The savings expressed in terms of ROM tons can becalculated according to the relation:

c2  = c1t2 (6)t1

Where

c2 = revised operating costc1 = original operating cost

t2 = throughput after pre-concentration

t1 = original ROM throughput

In the case of underground pre-concentration savings of up to 50% are also enjoyed in the area of backfill costs

through the generation of large quantities of cheap, competent backfill close to the active stopes. Crushing andconveying costs indicated above are replaced by the additional operating costs of pre-concentration. Substituting

the tonnage values indicated above into equation 6, the reduction in operating costs for the drift-and-fill method

costs detailed above can be calculated, and are shown in Table 7 below.

Table 7 – Mining and Milling Costs for Medium-Depth Cut-and-Fill with Pre-concentration

Mining

Mucking/

haulage Backfill

Pre-

concentration Hoisting

Surface

Haul Milling Tailings Total

46.90 8.19 12.14 2.08 1.33 3.11 5.13 1.4 80.28

Downstream operating cost savings for drift-and-fill method are of the order of 20% overall. Similar capital and

operating cost savings have been documented for the implementation of pre-concentration on surface. Capital

and operating cost savings as well as an increase in the effective capacity of the surface mill were observed(McCullough et al, 1999). The capital and operating costs savings precipitated through pre-concentration on

surface can only be maximised by implementing the concept underground. This observation has been borne out

in a number of case studies in the impact of implementing underground pre-concentration on cut-and-fill anddrift-and-fill mining scenarios at UBC (Schindler, 2001; Bamber, 2004).

4.  Conclusions

The basic economics of facilitating high-productivity mining methods in narrow-vein situations have been

 presented. Larger face dimensions and longer drill rounds can result in an unacceptable drop in the grade of the

ore in the panel. The adoption of an underground pre-concentration system with good metallurgical performancehas been shown to increase the grade of a round thus produced to a value superior to that of the original drift-

and-fill stope through the rejection of both external and interstitial waste. Simulation of the proposed integrated

mining and processing system has shown the physical possibility of adopting this method, and both literature

and calculation shows that there are capital and operating cost benefits to the system.

A case study focusing on the Onaping Depth orebody in Sudbury, Ontario is discussed. Mining of the orebody is

 presently considered unfeasible due to difficult ground conditions and associated safety risks resulting in

marginal economics. The implementation of underground pre-concentration at Onaping Depth has beensuggested: capital costs for downstream facilities such as hoisting and milling are increased by 6%, however

operating costs are reduced by between 20 and 40%. There is also a positive benefit to the project through an

improvement in ground control via the generation of substantial quantities of cheap, competent cemented backfill which can be pumped to the fill stopes as required.

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