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TABLE OF CONTENTS

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General Discussion • • • · • • • • • • • • • · 23

Economic Analysis of Present and ProposedMining Systems • • • • • • • • • • • • • • 26

Other Considerations • • • • • • • • • • • • • 27

APPROVAL ••••••

ACKNOWLEDGEMEN'lS •

REFERENCES •••

GENERAL CONCLUSIONS • • • • • • • • • • • • • • •• 32

INVESTIGATION OF THE ECONOMICS M~D PRACTICllliAPPLICATION OF BLOCK CAVING • • • • • • • • •

INTRODUCTION

LABORATORY INVESTIGATION •

CONCLUSIONS FROM LABORATORY IJIiVESTIGATION ••

GEOLOGY

HISTORY

SUI,IIMARY ••••

1

SUMMJI.RY

This thesis describes' an investigation and presents exper­

imental and economic data designed to ascertain if bulk mining methods

of a modified block caving type are economically feasible and applicable

to the iron deposits on the Gogebic Range. A study of block caving

mining methods, by use of mine models in the Lavorator-Ies of the University

of Wisconsin, personal visits to iron mines on the Gogebic range with

collaboration and consultation with mine operators, suggests that this

mining method can be applied successfully to the larger type iron

deposits of the district. The results of the investigation indicate

that increasing the present sub-LevaL Lnterval, from 75 to 150 feet is

practical and a substantial saving to mine operators is possible.

Care must be taken to select orebodies, or portions of them

having sufficient size and Uniformity of shape in order that the caving

operation may be successful. Attention is drawn to certain limitations

which should be carefully considered and special emphasis given to draw

control.

Good draw control will give high extraction and keep dilution

a t a minimum but it must be realized that extraction percentages and.

percent dilution figures are only as good as the original estimates of

ore tonnage. If the original ore tonnages estimates are in error, the

percent extraction will be erroneous. For this reason it is necessary

that the orebody be clearly outlined and any waste areas within the ore­

body known. If the inclusion of waste areas lowers the grade of the ore

too much to be marketable, then it will be necessary to mine these areas

.. . . '" .

by conventional sub-level caving and leave the waste area in place.

A comparison is made in this presentation of the economics

of sub-level caving with a 75 foot interval between subs and a block

caving system with a level interval of 150 feet. The basis for costs

is that of labor expense in terms of feet of advame of development

workings, feet of hole drilled in long hole drilling, and tons of ore

drawn during production per man shift worked.

Because it is not anticipated that changes in equipment will

be necessary for a block caving system, it is contended that an

analysis of cost based on labor required will give an adequate com­

parison between the two methods.

This method of estimating has the added advantage of being

unaffected by inflation or deflation since current monetary val.ues

may be substituted for manshifts and for tons of ore.

Footwall ore bodies, in the Plymouth formation of the Gogebic

are normally the largest and most uniform and will most readily lend

themselves to block caving and consequently this report limits its

investigation to. this larger type ore body.

2

3

INTRODUCTION

The Gogebic range, direct shipping iron ore center, extends

for about 70 miles between the v:Lllages of Mellen, Wisconsin and

Wakefield, lvu.chigan. (Fig.l) Two railroads, '~he Chicago & Northwestern

and the Minneapolis, St. Paul & Saul Ste. Marie, serve the district

and haul ore to their docks on Lake Superior at Ashland, Wisconsin.

From there the ore is taken by boat to the Lower lake ports. (1) The

climate of the region limits shipping to 7 or 8 months of the year.

Ore is stockpiled at the mines in the winter.

Rising costs of mining underground deposits of iron ore on

the Gogebic Iron Range present a problem for the mining engineer which

must be solved to assure the continuation of mining operations in the

area. High Ia bor , equipment and material expenses increase the cost

per ton for extraction of the ore from the ground. As mining progresses

downward, the greater expense of hoisting from lower elevations raises

its cost considerably. Supnort of underground openings become more

difficult at lower levels, due, in part, to the unstable condition of

the ground created by mining out the deposit above. The pumping of

water from the mines is also costly under the high pressure heads en­

countered at lower levels which may be as much as 4400 feet below the

collar of the shaft.

These higher expenses of operation alone are enough to make

the engineer look for a less expensive method of mining and in addition

there is the problem of marketing the product in competition with high-

•.'~.

4~

44

4 Z IW II! 2 , 4

Scale95 1 T r

Mil••

s 6

50

4~

48

47

46

4$

Figure I. - Loeetlon of the Gogeble Iron range.

4

grade taconite pellets produced in large tonnage operations in Minnesota

and Canada. Imports of high-grade direct shipping ore from Sweden,

Africa and South America are on the increase. 'Ihe taconite ore and the

majority of the ore imported from foreign lands comes from high tonnage,

low cost, open pit operations.

This increase in ccmpe td, tion from new sources of iron ore

ma~s it imperative that a bulk mining system be introduced on the Gogebic

Iron Range which will lower the unit cost per ton for extracting the ore

from the ground. Since the mining system now in use is a sublevel

caving mining method, it seems logical to consider increasing the dis­

tance between sub-levels and the drawing off of a higher back of ore

through each set of workings. To a certain extent this has already

been done. When sub-level caving was started on the Gogebic Range the

sub-interval was approximately 25 feet floor to floor. It has been in­

creased through the years to a maximum of 75 feet. Intervals of 40 to

50 feet have definitely been proved workable. The 75 foot interval is

still, to a certain degree, in the experimental stages but all evidence

seems to indicate it will be successful. (8)

Because of the success of increasing the distance between sub­

levels beyond 25 feet and the lower costs per ton experienced by so

doing, this study investigates the possibility of increasing the interval

between sublevels to 150 or possibly 200 feet. Since this Y()uld, in

many cases, involve the caving of the ore from the footwall to the hanging

wf,ll in one lift, the mining system is no longer a sublevel caving

operation but becomes a block caving mining system.

5

An:y bulk system which might be employed on the Gogebic Range

would, more than likely, give a lower grade product than that which is

mined by present methods. Because of the greater number of development

openings demanded by the present system, the orebody is well defined

and it is sometimes possible to leave low grade areas of waste rock in

place. A block caving system will have to include these low-grade areas

within the ore body and this will naturally lower the over-all grade of

the ore recovered. Because of this, block caving can only be applied

to the larger, more uniform deposits of the Gogebic Range. Small,

irregular deposits must still be mined by conventional mining methods.

The percentage of recovery of ore mined by block caving should

be relatively high. In districts where this system is used, recoverys of

90 to near 100 percent are common. (6)

HISTORY

The earliest trace of white man in the region was the presence

of two Frenchmen engaged in fur trading with the Indians at the present

site of Ashlani in 1658. Barnes ani Whitney, geologists, did the first

mapping of the area for Michigan in 1847 but failed to note the iron

bearing formations. (1) Iron bearing structure was first noted in 1848

by A. Raniall but it wasn't until 1884 that the first shipment (1,022

tons) of iron ore was made from the Colby Mine near Bessemer, l\Iichigan.

Since the first shipment, and through 1956, a total of 301,428,189

tons have been shipped.(2)

The range has a total length of approxima tely 70 miles with

25 miles in Michigan ani the remainder in Wisconsin. However, a portion

of about 12 miles in Michigan and 4 miles in Wisconsin has produced all

the marketable iron ore mined on the Gogebic Range except for about

4,000,000 tons. (3)

Early mining was by primitive open pit methods and by the

sinking of shallow shafts under-ground by hand labor and hand operated

hoists. In 1886 and 1887, speculative capital was brought into the

region to the extent of $1,000,000,000 and mining operations were

expanded tremendously. (3)

An inevitable cutback in operations in late 1887 took the

savings of many of the small investors aid many mines were closed.

The larger companies weathered the storm and, in spite of the speculative

failure, ore production rose on the range. Production from the Gogebic

6

7

range has been relatively constant since, except during periods of

depression. Production during recent years has averaged about 5 per

cent of the total production in the Lake Superior District which in

1956 totaled 81,913,815 tons.

Top slicing and sub-level caving were the mining systems Which

were the predomina te mining methods employed on this range. There was

and still is some sub-level stoping in sections of the ore body which

are strong enough to stand during mining but most of the ore produced

on the Gogebic has come from sub-level caving or some modification

thereof. The trend in recent years has been to induce breakage of the

ore by longhole drilling and increase production by extending the dis­

tance between sub-levels.

The main producing companies on the range at the present time

are Pf.ckands Maither Company, Oglebay Norton Mining Company and North

Range Mining Co. Pi.ckands Mather Company is the largest of the three

and is an operating company. It does not own property but mines the

ore for the owners on a l~ase or royalty basis. Mines operated by

Pickands Mather Company and subsidiaries include the Peterson, Sunday'

Lake, Geneva, and Newport Mines in Michigan and the Cary Mine in

Wisconsin. Oglebay Norton Mining Co. operates the Montreal Mine in

Montreal, Wisconsin and North Range Mining Company operates the Penokee

Mine in Ironwood, Michigan.

Table 1 shows the tonnage and average analyses of Gogebic

iron ore shipments for 1955 and the average tonnages and analyses

for the past ten-year period.

TABLE I

ORE GOGEBIC RANGEGross Tons Percent Percent Percent Percent Percent

Shipped PerCent FeNat. Dry Phos Dry Si02 Dry lIIln Moist.

Bessemer 49,001 1.0 52.33 .042 9.56 .33 10.78

Low PhoseNon-Bessemer 4,837,553 99.0 52.49 .077 9.35 .48 10.95

Total 4,886,554 100.0 52.48 .076 9.36 .48 10.95

Average1946-'55 4,665,104 52.49 .076 8.78 .56 11.41

From Bulletin of the University of Minnesota, Mining Directory Issue

by Henry H. Wade and Mildred R. Alm.

GEOLOGY

Ore has been found largely as massive bodies of hematite

along troughs formed by intrusive dikes that intersect ste<:lply dip­

ping beds of sedimentaries. The range .Ls a sedimentary series of

ferruginous chert known as the Ironwood formation. The formation is

500 to 600 feet in thickness and the beds dip 60 to 70 degrees north

with a general strike for the range of approximately N.63° E.(Fig.2)

In general, the ore is related to a longitudinal or bedding

fault that divides the Ironwood formation into footwall and hanging

wall zones. Ore may be found eitherctbdve or below this fault. The

early explorations first found ore bodies lying on the quartzite

footwall above pitching dikes, but later development demonstrated that

the ore bodies Were distributed throughout the Ironwood formation

from footwall to hangingwall. Transverse faulting, which has shattered

the formation, is recognized as a favorable factdrin allowing for the

ciroulation of ground water which leached the silica from the formations

leaVing the iron enriched deposits lying upon the cross dikes.(l)

The beds within the Erorrso od formation have certain distin­

guishing charactaristics throughout the length of the range, and upon

these charactaristics the following subdivisions have been made from

the top downward:

1. Pabst member, cherty and fragmental, and ferruginous

slate beds.

GOGES IC RANGEGEN ERALI ZED CROSS SECTION

LOOK ING WEST

\IRON FORMATION

Figure 2

2. Anvil, ferruginous chert member.

3. Pence, ferrugenous slate member.

4. Norrie, ferruginous chert member.

5. Yale, interbedded chert and ferruginous slate.

6. Plymouth, fenmginous chert member.

10

LABORATORY INVESTIGATION

Equipment for experimentaJ. work on this project consisted

of two mine modeLs which were designed to demonstrate fundamentals;

the acceptance of which seems necessary if block caving is to be

successful. A view of the first model is shown in Fig. 3.

This mine IIIOdel was designed at a scale· of 1 inch equals

6 1/2 feet. HaJ.f inch plywood was used for construction of the back

and sides and a double thickness window glass was used in the front

so that the action of the draw could be observed. MetaJ. funnels with

a half inch diameter discharge were placed along the bottom and sides

of the model as draw points. The top of the model was open so that

sand, crushed limestone or iron ore could be placed in it for various

experiments. Rubber stoppers were used in the funnel openings so that

control of draw of the material was assured.

Essentially, the model represents a section taken perpendic­

ular to the strike of the deposit, looking west. The left hand side

of the model representing the quart~ite footwall and the bottom the

horizontaJ. plane of a system of development workings. Draw points

were put in to represent mill holes from development slice drifts

along the footwall and in the horizontal plane at the bottom. Each

fUnnel would influence an area, according· to the scale of 1 inch

equals 6 1/2 feet, of approximately 16 by 19 feet or approximately

300 square feet.

. .

Jol

Fig. 3

Plywood Sides 7 ~

......"l\lf\J

T

Mine lb de l No. 1

Glass Front

--------- 29 "-----~

FunnelDraw Poi n ts

12

Expe riment No . 1 - Model No. 1

Ini t ial experimental work was done with a f oundry s an d which

was dry an d flowed ve ry eas ily . Angle of repose of the sand was ap­

proxima t ely 30 de grees from the hori zontal. The s an d was placed in

the modeL and the No . 3 draw poin t from the l eft on t he bottom was

opened.

Result:

The aa rd flowed freely from the ope ning but the r e was no

movement of sand particle s evide nt behi nd the glas s. Movement took

place on the top of the s and and a conical depression was noted

directly above the draw point. A piece of gr avel was dropped into

this and was very shortly drawn off at the draw point indicating

there was a r apidly developed, narrow, pipe-l ike channel of f'Lowdng

mate r i al di rectly abo ve the opening. The s ize of this channel mus t

not have been l arge s ince no movemant could be noted in the sand behind

the glas s an d t he model was only 2 1/2 i nche s wide . Mat er i al was being

drawn off mainly from the top of the mass as Vias evidenced by the r apid

pa ssage of the gr a ve l through the s and , by t he cone tha t deve loped at

the top of it and by t he l ack of movement in the s and behind the glas s .

Conclusion:

It was concluded from much repet i t i on of this f irs t exper ime nt

that the angle of draw ( angle of di rection of movement of material with

the horizontal) in caving very fine mate rial appr oximat es 90 degrees and

the f low pattern is mainly tubular above the opening. It was decided

13

that a coarser product should be tried in order to observe what effect

this would have on the angle of draw.

Experiment No. 2 - Model No.1

Limestone was crushed to 1/4 inch size in the laboratory

crushers. The fines created in the crushing operation were 'left in

the product so that it ranged in size from 1/4 inch down to dust.

The model was filled with the crushed limestone and "number 3 draw

point was opened. The material did not flow readily and it was neces­

sary to poke at the hole with a wire to keep it flowing.

Result:

No movement was noted in the limestone near the bottom of

the mass but near the top some movement could be seen behind the glass.

Coning took place at the top of the limestone similar to that Which

occured with the sand in Experiment 1 but it was not as pronounced.

The angle of movement which showed through the glass was measured at

about 86 degrees with the horizontal which indicates that the angle

of draw was again nearly vertical. No. 3 draw point was closed and

various other draw points were opened and closed to observe the draw

through the glass. It was found that if considerable material were

drawn from an opening and then the point next to it opened, there was

a tendency for the second to draw material from an area above the first.

This suggested that a vertical plane of weakness was created above the

first point drawn and when the second point was pulled it broke into

this weakened channel above the first and drew material from it. Il­

lustrated in Fig. 4. This was the only case in which the angle of draw

was anything but vertical during the tests conducted.

Conclusion:

It was concluded that only small amounts of material should

be drawn from anyone finger when it is opened to prevent the formation

of a draw channel to the top through the mass of ore. In practice,

overdrawing of one finger 'AOuld create a channel which naturally would

ultimately allow entrance of capping materials into the orebody and

cause exessive dilution. It follows that all fingers should be drawn

uniformly to elimin~e the development of the aforementioned planes

of weakness Which, in a given opening will cause the draw of material

from one side or the other when or if the material over the first draw

point is strong enough to be self-supporting. It was therefore decided

to use crushed iron ore in the model and to layer material on the top of

the iron ore so that the best system of drawing the fingers could be

ascertained.

Experiment No. 3 - Model No.1

Iron ore from the Geneva Mine, Ironwood, Michigan was crushed

to 1/4 inch in the labora tory crushers and the fines were left in the

ore, giving a broad size range product. The ore was placed in the model

and a 2 inch layer of grinding pebbels approximately 1/2 inch in diameter

was placed on the top to Fingers were drawn from

left to right and only ore were taken out of the opening

Stone

Crushed

Break Through

into Channel

Fig . 4

Plan of Drawpoints

Cross- section of Model

. -

15

during each draw period. The drawing process was r epeated across

t he model agai n and agai n until appr oximat e l y 80 percent of t he are

vias drawn out.

Results:

It was found t ha t even drawing of openings tended to keep

the line between the pebbles and ar e rela tively uniform but t he re was

still a tendency f or t he channels t o develop above the openings. The

capping moved down through t he or e , di luting it and c l os i ng of f the

draw point ope ni ngs before all of i t was drawn of f .

Conclusions:

It was s us pected that due to the narrowness of the model

(2 1/2 i nches) and the l ack of weight on top of t he are t ha t be tween

dr aw poi nts t he are was, t o a cer tain extent, self supporting al l owi ng

the channels to devel op ai d causing t he dilution by t he pe bbles or

ca ppfng , To correct this situation and t o more nearly.simulate the

conditions f ound in ac t ual practice, it was decided to construct

another model which would exert a force on the top of the are normally

c reated by the weight of the capping a nd have a grea ter l ateral extent

so that the ore would not tend to support itself.

1bdel No . 2,

A view of the s econd model is shown in Fig. 5. This model

was designed at a scale of 1 inch equals 10 f eet . The sides were made

by placing t hree foundry f lasks on top of each o t her and the bottom

was a 1/4 inch aluminwn plate with 1/2 inch holes drilled in it f or

Air Intake

Anchor Bolt

Aluminum Head

dry Flask

; - - - - -..!.!.-I ' \. ,, ,\ I

I{' \\1Diaphram

I I I I' II -- - - - - - _/

III

I

Fig. 5

u U u

Wood Stand

/ - - -I- -- _'-!._( ,I ,

I

Mine Model No. 2

r - r lI I I

I . I II I Diaphram J I\ J"'-- - - - - - ..... I

IIIII

LJ

- - -/6$."

dn n

1" X 1" x 1 8" Angle

@--

l.v l.y... Foun

0 ,,"-~

~V

-- - '--i!=

II.III

16

draw poi nt s on 2 inch centers. The sides and bottom were held rigidly

together by 1/4 inch bolts and screws. The top consisted of a cast

aluminum he ad to which a rubber diaphram was f as tened. An a i r pipe

was t apped into t he he ad so that air c oul d be i njec ted into it to ex­

pand the diaphram to the f ull s:4,e of t he node l , Whe n the model was

in use the head was held t o the supoor t i ng s i de s by a 2 1/4 inch b ol t

and a I " x I " X 1/8" angle iron across t he top .

Dur i ng operati on , the hol es in t he bottom were plugged with

r ubber stoppers and t he top unit was renoved , Sand, crushed iron or e

and l i mes tone were re~pectively pu t i n t he f l asks to a depth of 10

i nches in each case . The di.aphr-am an d head uni, t was replaced and

bolted down so that a ir could be i nj ected i nto t he df.aphram a t any

des i r ed pressure. An air pressure s ource of 90 p .s .i . gauge was avail­

able for t h i s e :cperiment . Between the source and the head an air reg­

ulator and a pr es s ure gauge were installed in t he l i ne so that a cons­

tant pressure could be maintained within the model at all times.

The purpose of the diaphram was to s in.ulate condit ions which

might be encountered when caving ore at a considerable dept h underground .

In caving operations, the capping material exe r t s a force upon the or e

as i t caves and, it is the opi ni on of t he au t hor , that in the case of

the iron ore of the Gogebic, t his force i s necessary an:! essential to

a block c aving ope r a t i on. The capping a i ds in br e aking up the or e and

breaking down the natural pressure a r ch which forms over any opening

i n the ground. I t is also conte nded that t he extra weight of the capping

will brea k down t he side wal l s of channel s which may form over draw

point s.

17

the ore. Material was drawn

1 inch equals 10 f'eetScale of' Model No. 2

Experiment No.1 Model No.2

The model was :filled with 1;4 inch limestone to a depth of'

1 cu. f't. of' ore weighs 200 Ibs.

Volume to be drawn 4 Sq. in. x 10 in. ~ 40 cu. inches

Since it is desired to duplicate conditions 1000 f'eet under-

an ore body lying under 1000 f'eet of' capping is caved. It is assumed

Vertical height is 10 inches

culations represent the analagous conditions USing laboratory model:

Ef'f'ective draw area of' each draw point equals 2;' x 2" or 4 sq. inches

develop, on a Labor-a tory scale, similar conditions encountered when

1728 cu. inches per cu. f'oot

the ore body is 100 f'eet in vertical thickness. The f'ollowing cal-

For purposes of' this experiment, it was decided to try to

200 Ibs. x 0.02307 cu. f't•• 4.614 Ibs. of' ore per draw point

40 cu. in.1728 cu. in•• 0.02307 cu. f't.

of' 100 f'eet of' ore.

ground, if' 1.153 is multiplied by 10 a pressure of' 11.53 Lbs, per square

4.614 Ibs.4 sq. in. • 1.153 Ibs. per sq. in. proportionally represents pressure

inch in the diaphram will proportionatly represent the desi,ed condition.

See Fig. 6.

10 inches. The diaphram head was bolted in place and an air pressure

of' 11.5 p.s.i. was exerted on

Pressure here equals

1 . 1.53 lbs per sq . i nch

o

Fig. 6

'--- 2 "----

... ,

J#

1/2 n Draw opening

U •.53 lbs per s q. i nch

Pressure equal to 100 inches of

cupping f rom diaphram

18

evenly from draw points back a nd forth across t he model. About 15

cubi c i nches was drawn from each po i nt as it was opened . After about

half the limestone was drawn from the model the pr essure was released

and t he top removed so that the contact between t he limestone and t he

diaphram could be observed. I t was no t always pos s i bl e to ge t the

desired amount, of material from each draw point due t o clogging and

in s ome cases too muc h was drawn because of f r ee running of the material.

Results :

The surface con tact of the l i me stone was concaved downwar d

and t he mat erial a t the s ides had not drawn as readily as the middle.

Conclusion:

Since the di.ap hr-an had to be fol ded slightly on the s ides

and t he s ides of the model t apered out wa rd t oward the bottom, it was

not une xp ec ted that t he surface of the limestone would be concaved t o

a certain extent. Ther e was a tendency to over dr aw in the middle due

to the ease with which t he material flowed out of draw poi n t s . Draw

points a t the sides and ends hung up quicker an.; draw was more difficult .

It was decided to use iron or e in the model and t o put a layer of grin­

ding pe bbles on the t op so that dilution conditi ons could be as cer t ai ned .

Experiment No . 2 Model No . 2

Crushed iron ore was pl ace d in the model to a de pth of 9

inches and a 1 i nch l aye r of quc.rtz grinding pebbles was pl aced on

top. The dfaphran he ad was pu t in pl ace a nd secured and air pressure

of 11.5 p.s.i. appl ied t o t he head . Or e was dral~ unif ormly ba ck a nd

19

forth across the model, at a rate of approximately 10 cubic inches

of material drawn at each opening.

Results:

No dilution was noted through the first two drawings of

all openings but on the third opening of points near the middle of

the modal , pebbles clogged the holes. Holes to the outside of the

model didn't become clogged with pebbles until after the third and

sometimes fourth drawing periods. It was determined that approximately

75 percent of the ore was drawn before dilution became serious. Again

when the model was opened there was a build up of materials on the

/!ides.

Conclusions:

It was concluded that too much ore was drawn at each drawing

period. Approximately 25 percent of the ore was drawn at each draw­

point each time the hole was opened. Drawing this rather high per­

centage of the total volume above a drawpoint each time it was opened

disrupted the division between the pebbles and the ore causing exces­

sive dilution. It was decided to run the sane experiment again but to

limit draw to 10 percent of the volume above the drawpoint at each

drawing period.

Experiment No. 3 Model No.2

The model was prepared in the same manner as was done in

Experiment 2. Approximately 4 cubic inches of ore was drawn at each

drawing period.

20

Results:

Extraction rate was very good in this experiment as approx­

imately 90 percent of the material was drawn out of the model before

serious dilution from the pebbles occured , The 10 percent of the ore

which was not drawn was on the tapered sides and in pillars between

openings in the base plate of the model.

Conclusions:

It would appear from the procedure that the uniform drawing

of a relatively small amount of ore results in a mimimwn of dilution

by the capping. In fact the amount of dilution is lessened to a degree

where its effect is not considered dangerous to the economy of the op­

eration. Keeping the draw uniform and drawing not more than 10 per­

cent of the volwne above each finger opening 'nill give optimum results.

It is not anticipated that decreasing the amount of the draw

to 10 percent will limit the production from a given area to below that

requi.red to maintain the normal capac i, ty requirements of the mine. If

the ore body is 100 feet thick then 10 percent of the draw would be

10 vertical feet of ore over an area of 400 square feet. Using a tonnage

factor of 200 lbs. per cubic foot this would amount to 200 tons which

could be drawn each time the finger was opened. A production drii't

would normally be worked having at least 4 finger openings so that

800 tons could be removed before it is necessary to move on to another

group of 4 fingers. This represents at least 5 shifts of production

if two men can draw 150 tons of ore in one shift.

00NCLUSIONS FROM LABORATORY INVESTIGATION

I. Dilution

Drawing of small amounts of ore from each point in rotation

will tend to keep dilution at a minimum and the entire mass of ore can

be lowered wi th the least amount of disturbance to the ore-capping line.

No mere than 10 percent of the ore vertically above a finger opening

should be drawn at any one time. In the case of the 100 feet high ore

body this means that the finger must be opened, pulled and closed about

10 times before it is empty and capping appears.

II. Angle of Draw

Experiment shows that for all practical purposes, the angle

of draw is approximately 90 degrees with the horizontal. Thus the

material drawn from a finger comes from an area directly above the

finger opening. The largest deviation from the 90 degree angle was

with the coarse material and the measured angle was 86 degrees. This

small difference is considered to be of little significance and prac­

tically the 90 degree angle of draw predominates. It is here worthy of

note that this conclusion supports the work done by other investigators

under analegous conditions. (5)

III. Channeling above Draw Points

Channeling will tend to develop to a greater extent in fine

material than in coarse. If this condition is encountered, the finger

opening should be closed to allow a pressure to build up on the side

walls of the channel so they will be crushed thereby eliminating the

21

22

the channel. Channeling is not serious until the channel reaches

the dividing line between the ore and the capping and then it allows

the capping to drop through into the ore and serious dilution results.

Pressure from over-lying ore and capping is desireable as it tends to

reduce channels in the ore.

INVESUGATIOH OF THE EC ON01;;ICS AND PRACTI CAL APPLI CATION OF BLOCK

CAVING

Gener al Discussion

Subleve l ca ving has been prac t iced on t he Gogebic r ange for

many years and generally is well organized a nd understood by miners

an d s upervisors. I f t he present system i s t o be changed there must

be considerable r e ason for doi ng so . I f a grea ter pr of i t can be ob­

tained by cha nging the present procedur e ser i ous conside r at i on should

be gi ve n any i deas put forth. As was s ta ted in t he Introduction,

ris ing c os t s of mining have influenced the engi neer to look for lower

cost methods.

Subl eve l caving has t he advantage of being a s afe mining

system. It Will, when properly conduc ted , yield a hi gh ext rac t i on

and produce a c l ean or e. Generally, sub-level ca ving is the i nte rmed ­

i ate caving me thod be t ween top-slicing and block caving . Thi s i s i l­

lustra ted in Table 2 , (1) which gi ves a compari s on of the t hree cavi ng

systems.

In order of merit of caving sys tems, bl ock caving r a nks

first as a che ap mining system, highest pe r cent of ar e won oy caving ,

low t i mber consumption , e ase of vent i l a tion and l arge output gained

f r om a gi ven a rea . Bl ock caving r an ks l as t in clean mining, per cen­

tage ex t r acti on , f l exibil ity, and control of cavi ng.

From t he previous paragr aph it i s i ndicated that if block

caving can be appl ied it will r e sult in substantial savings i n mining

2)

be well controlled.

rs

BC

TS

TS

Be

of merit3TS

BJ

SC TS

se

SC

se

so

SC

se

SC

SC

TS : Top slicing.

Usual order2

SC

SC

Be

TS

BJ

TS

TS

1'S

1

BC

a

TABLE 2

COMPlillISON OF CAVING METHODS (4)

Natural ventilation

Large output from given area

BJ: Block caving SC: Sub level caving

24

Percent of ore won by caving

costs. The question arises as to whether or not the disadvantages

Control of caving

mentioned laboratory research, and if properly done, a clean product

a Fire hazard varies directly as timber consumption.

From Mining Engineers' Handbook by Robert Peele - Vol. I

All of the disadvantages can be controlled if proper planning

ani supervision are given. Close control of draw at finger openings

Chance of losing ore

Percentage extraction

Timber consumption

of the system can be overcome.

can be obtained, a high extraction realized and the caving action can

Cheap mining costs

Close grading of are

is essential in any caving system as has been supported by the afore-

Clean mining

Flexibility

From the standpoint of

25

Lack of flexibility of the system should not prove detrimental

if ore bodies which are to be caved are sui table and are selected with

foresight. There should be no reason for changing the system once it

is started if the ore body is of a cavable nature and well-defined.

Since sub-level caving has been used for years it is defini tely established

tha t the ground is cavab.Le ,

As to the chance of losing some of the ore, this ought to be

anticipated if block caving is to be used. The possibility of losing

ore can be greatly lessened by good drilling practices to outline the

ore, by close draw control and by mining of sections of the ore body

which can not be caved by some other system. Lower unit mining costs

should offset any losses of ore due to the bulk caving system.

The larger type, footwall deposits on the Gogebic range

would seem to lend themselves to block caving. One deposit may be up

to 120 feet Wide, 200 feet high and run several hundreds of feet along

its strike. The only serious disadvantage in the deposit, as far as

block caving is concerned, is the fact that the footwall is inclined

at approximately 65 to 70 degrees. This disadvantage can be overcome

by undercutting the footwall side of the ore body first and inducing

the caving of the footwall section of the ore body previous to caving

of the main section. Induoed saving can be brought about by drilling

long drill holes with peroussion machines from drifts in the ore or

footwall, blasting the area of the holes above the undercut at the foot­

wall and drawing off the ore at draw points. This is actually a mod-

26

Hied shrinkage stope mihingmethod. It may be found that it is not

necessary to do a great deal of drilling to get the ground to cave if

the ground is of a particularly weak nature.

Economic Analysis of Present and Proposed Mining Systems

The typical ore body to be investigated is 400 feet by 120

feet in plan and the level interval is 150 feet. 'l'he footwall ai d

hanging wall of the deposit are approximately parallel and dip at 65

degrees. Fig. 7 is a cross-section of the orebody looking west and

Fig. 8 is a plan view of the deposit in the horizontal plane of the

sub-level.

Table 3 is a comparison of manshifts required for mining the

deposit in the currently adopted 75 foot intervals and the proposed

150 foot intervals.

Fig. 9 am 10 show the layout of workings for the proposed

system which caves 150 feet of ore or the entire orebody between levels.

Table 3 shows a definite saving to the mine operators of

4532 manshifts if this deposit is mined using the 150 foot interval.

Other savings that will result which are not evident from this analysis

are in moving equipment from one working place to amther-, loss of time

in starting new development headings, easier supervision because of

fewer working places and small labor force necessary to obtain the

required production.

TABLE 3

ANALYSIS OF MINING SYSTEMS BASED ON CURRENT COSTS _ 1958

*Longho1e Drilling from Drilling Subs - 10 Foot Interval Between Stations]60 Feet of Drilling Per StationTime to Set Up and Drill Holes = 2 SlJifts or 4 Manshitts

*Millho1es - Includes Longho1e DrilJ,ing aIldNecessary Raising to Open Millho1es1200 Feet of Longho.Ie Drilling a. 200 feet per shift = 12 Manshifts

12 Feet of Mill Raising a 8 feet per shift = 3 MansniftsTotal 15 Manshifts

MAN....SHIFTSSAVED

150 Foot IntervalTOTAL MAN- TOTAL

LENG'IH LUTH OR SHIFTS MAN-UNITS PER FT. SHIFTS--•

2400 • 6 200 1200 1.0 1200 1200800 • 1 400 400 1.0 400 400410 • 2 120 240 1.0 240 170260 • 2 70 140 1.0 140 120540 • 2 165 330 1.0 330 210300 • 1 75 75 1.0 75 225495 • 1 165 165 :1..5 248 247765 • 1 850 850 0.9 765

]600 .120 -- 120 15 1800 1800--

•• 720,000 Tons to Draw-

14440 > 50 Tons/Mansmft 14440,•• 40 Stations a 4 Mansmfts/

]20 · . Station 160 160•• 2 Men/Shift During 2400

4800 • Drawing Shifts 4800•

29130 • 24598 453224.7. 720,00OTons~24.598

Manshifts = 29.3

1.01.01.01.01.01.01.50.9

15

>

2400800410260540300330850240

TOTAL MAN- TOTAL •LGTH OR SHIFTS MAN- .NO.

UNITS PER FT.. SHIF1'S.

720,000 Tons to Draw­50 Ton/Manshift

80 Stations a 4 Manshifts/Station

2 Men/Shift During 2400Drawing Shifts

2 75 Foot Intervals

Slushing Drifts 12 200Drilling Subs 2 400X-Cuts in Ore 2 205X-Cuts in Rock 2 130Double Box Raise in Ore 6 90Branch Raise in Ore 6 50Pemble Box Raise in Rock 2 165x-Cut Haulage Drift 1 850J4j.ll Holes * 240

WORK NO.. LENGTH

Longhole Drillingll-

prawing Ore

Timb.er Repair

TotalTon Per Manshift 720,000 Tons~29,130 Manshifts =

rr:,'-

Present Mi ning System

Ore Pas s

Cross-section - Looking We s tI

Scale 1"-30'

• • • 0 • •

- .. :...:: -= ~ . : .:-':

... . .. , . ., '.

a- Drill Sub

:. -.: ; -= -:. ..

Area to be Caved

..-

~\\:~~~hOUt_~'i~\~

. . --":'":

Sub-l evel

~ .. : ,.

_:::_.;':_= : :0_ : - -

-- -- -- _ _ - _ » .u.J (.I,u·~...un U~:.L.J. ~ " i I\ J \ \ )

_ 2nd Level_

"

~

"~

1st Level

\\- -\ \\ \\\\ \

\\ \ \." }'"( / ' .\ I ( ./' .. \ )"l I , . . ,

" ~ 4tV ~ •

~i \ ' Y Y \I V "~ !~) \ , . o j I ... _ _ i I 4 . . . . _

" \\\ vY" VentUa tion &<

'\ SUpply Raise

\\\ \\\\\ II \11 I I

• ( J ' I ,

'>:I "....Oll•....

I

Plan at Sub-Level Elevation - Scale 1" • 50'

YaleHanging Wall

.. . .. .. .. \ ..... .., .... .. .. .. .. , .. "..

. ..... .. '\

~ .. . .... --- , _. """

--c Em of Block Em of Block ~I ' I

' .. P"-uth . .

I n' ~- :n' II . . . " . . . . . .'. - . .~ ~ - . ' . . . - .. . . . - . - ... . . . " " • - -. ~ '~' , ' ., ., , , ., . ' . . . .

: [ \ - - -.. .. I

f '-:'- ::::::-- - - -;,I~~:-~ ·~· :: ' :..::.:::: :: _..~ ~ ,...= ==-~~_.- =-=- -=--=--=--_-.Jlr1J 1 j ll&-Sub.- -= - '- -- - - - -- -. . ' - . . -.- - , ' .._ ' - " " . , ..: ,-,-, .-.-. ,- ,-. -. -, .-.-..;-; ,~ I'J iJ Iil

00

"':l...""•

.. .. . .. .. . .. . .. .. . -- .. ..... , • • • • • • , . ' I , • •• • • • • '1

Footwall

Ventilation & SupplyRaise

• , , • • r , ,. , r -· • i - i -,-, •

Scale I n. 30 '

Proposed Mining Sys tem

__ --, Cross - seotion - Looking We s t

-~

Caving Block

Drill Sub

'.\~ Floor Pillar--_\ \. _ Will Cave\ \

\ \

\\ Footwall _ .-\.. _. _

Rai ses in Ore - - \

\\ Each End of Block Holes Dr illed

\Here Where

\ Needed

\\ \ 'rY-\ \

~p=~~~%~~~Jj~l_

\ \ Ventilati on\r Raise

\ \\ \\ \

\\ -\; -- -- - -- - -- +-.'\--};;;{.-.!

\- 71-- - --, -.....L..-.J. 2nd Level

~~

'" "

'%J,...-(JQ

•

..

I

I.,1._ ." , I • , • •

loading X- Cut

, ,

"" "' 1 ",

mray Raise

. . . , ' . " ... ."•,

t • •• , •• • • · · ,··

.-'. • I • • •

•,' .. , . .. ,' .. .. ' ..

Ventilation & Supply Raiseto 1s t Level

.. ... .. ...

Haulage X-Cut

. .. .. .. .. .. . ..

.. , . . .. .. . .. . .. . .. .. .. .Breaking Raise

..' .. . .. . . .. '" .. ..

Loading X-Cut

.. .. ....

Plan at 2m. Level Elevation - Scale 1" • 50'

• •• I • • • .. • • .. • •

... . ..

..

...... .......... ..

Mamray Raise

Yale

Plymouth

.. ... .... .. .. .. ......

I="' .... ......-" :I I:·· ···· ·..· ·· .. 'll:'.-........... --;I. .: . - .. - - - ..

/······ ·.. · ·.. ·11····· .. ···· ··'1 1··· ·· .. · ··I.. .. .. .. . ... . .. . .. .. ...... __ .

b

"l...Ill>•

27

Other Considerations

I. Timbering

A r eappr aisal of timbering pr ac t i ce s on the Gogebic r ange

will be neces s ary. Present timber as it i s i nstal led needs only to

sta ni while 25 to 75 f eet of ore i s drawn off t hrough the wor kings .

Timber now used stands up under pr esent conditions but Wit h block

caving it will have to withstand great er pressur e and mor e wear from

the ore as thicknesses of 150 or more feet of ore a re drawn t hrough

one level of workings.

Round timber, 11 to 15 inches in diameter, used for sets at

the present time are placed on 4 to 5 foot centers. Split cedar lag­

gi ng is used to block out the se t s and little or no e xtra r einforcement

is put in a t finger openings. If similar ti l: ,ber is to be used in a

block caving oper ation the dis t ance between sets must, be reduced to

'P pr oximatel y 2 to 2 1/2 fee t a nd reinforcement installed a t f inger

openings. This can be done by installing a pony set inside t he finger

behind t he drift set and by placing l a r ge, heavy poles across drift

set caps a t finger openings.

Steel sets of the yielding arch and r igid ar ch type a r e used

in some places within the mines a t the present t i lle , usually in areas

where grea ter weights are ant i c ipated . If the weigh ts encountered are

too grea t and the s e ts fail, it i s usually the l e g of the s et which

bends. This is true in the 3 piece yielding arch set. It is t he

author's opinion that f ailure resul ts from excessive side pre s sure i n

the drift. The cap, being able to slip in the cl.amps a t the top of

28

the legs can relieve i tBeIf of the pressure and retain its original

strength. The base of the set is held rigid and cannot move so that

failure takes place in the leg. Because of this condition it is

recommended that, in areas of unstable ground, where large pressures

are expected, full circle yieldable arch sets be used to support the

openings. In full oircle yielding steel ,sets the sets can yield Without

failure regardless of the direction of pressure.

It is difficult to estimate exactly the increase in cost which

will result from changing the present timbering practice. Because of the

increased level interval in the proposed system approximately twice as

much ore will be recovered for each foot of development opening driven

azd supported, thus lowering material cost. In the analysis on Table 3

it may be noted that 4800 manshifts were allowed for timber repair in

both systems, consequently there are twice as many manshifts allowed per

foot for timber repair under the new system than under the present method.

It i6_ the author's opinion that the aforementioned factors will offset, ,

any additional costs resulting from necessary changes in timbering

practices.

II. Explorationary Drilling

In the present system of mining, the sub-levels open up the

ore body in many places am serve to explore the ore body so that a

minimum of explorationary drilling is necessary. In the proposed system,

it will be nedessary to drillDlore holes to olitline the ore body as it

will be explored on only one level by the workings" If ElXplora tory

drill holes are drilled at 100 foot intervals along the strike of the

29

of the deposit, it is more than likely that 3 or 4 holes from each

drill station will outline the deposit. These holes will probably

average about 175 feet in length. Samples of material drilled should

be taken each five feet of hole and assayed. The drilling program

should be under the direction of the mine engineer so that he can

stay ahead of the drilling program and layout holes at the most ad­

vantageous angle to obtain the greatest amount of inforlll!l.tion. The

presently used Gardener-Denver 123-4 1/2 inch drifter machine will drill

these holes so that it is not necessary to purchase new equipment for

the drilling program and the holes are not long so this cost is not con­

sidered in this report.

III. Draw Control

Operators in the district have indicated a concern as to

the action of the yale Slates in the hanging wall, during caving

operations. Normally, the slates are weak and break into fine material

when caved, thus creating a possible dilution problem. The ores of the

Gogebic tend to break to a relatively fine size when caved. Consequent­

ly, the degree of penetration of the ore by the slates can be expected

to be relatively slight, due to the greater hindered settling effect of

-this ore as compared to coarse breaking mater-La'I ,

As has been formerly stated, draw control must be accurately

and scientifically controlled if block caving is to be successful and

ore must be drawn evenly to maintain a line between the capping and

ore and prevent uncontrolled dilution. of' the. product., Therefore the

30

schedule of draw should be established by the engineering department,

which keeps all records on draw control. One responsible individual

within that department should be in charge of all scheduling and see

that the operators follow engineering specifications.

Shift foremen should be given draw control sheets at the

start of the shift for each drift from which production will be ob­

tained. These sheets should have the finger nwnbers on them and how

many cars, scrapers or tons should be drawn from each. The shift

foreman should have his men follow these draw sheets as closely as

possible and have them record the actual amount of ore taken from each

finger. Comments should be noted on the draw sheet which may be of

value to the draw control engineer. The sheets should be turned in to

the engineering department at the end of the shift for recording.

It may be necessary to start a school or lecture series to

educa te the foremen and miners as to the importance of accurate recor­

ding of draw on the sheets. Models with glass fronts in which the

miners can see the effects of uneven draw practices would be especially

effective in driving home the need for accurate records.

Graphs should be prepared by the engineering department

showing the draw from the finger openings as compared with the estimated

total tonage to be drawn. The graphs should be prepared to scale on

both north-south and E;last"'west sections and be brought up to date

weekly. It will be obvious from the graphs when fingers are being

drawn too fast or too slowly and then corrections for erroneous draw

can be made.

31

Repair work in drifts, behavior of oertain fingers, grade of

ore drawn, oontrol of dilution, and pressure exerted on drawing seotions

are faotors affeoting drawing. Praotioal exparienoe and olose obser­

vation are important requisites forsucoess of the operation and for

this reason the draw control engineer should be in olose contact with

oonditions in drawing areas so he can adjust the draw sheets aocording­

ly. (1)

rv. Size of Caving Block

It is difficult to predict exaotly what size blook is the

minimum that can be caved in a new caving projeot. Based upon the

exper-Ience of other mines wlilich use block cavfng , it would seem that

an area 120 feet by 80 feet could be caved readily.

At the Greater Butte Project in Butte, Montana, blooks are

limited in size to 80 feet along the strike and 120 to 150 feet across

the deposit. (6) At the Sunrise Mine, Platte County, Wyoming, panels

100 feet by 90 to 120 feet are oonsidered to be the best size for

oaving.(7)

For purposes of analyses, a blook 400 feet along the strike

and 120 feet wide was oonsidered. It is the author's opinion that if

a blook 200 feet along the strike by 120 feet wide were tried it would

prove successful.

GEiITilltAb CONCLUSIONS

It is concluded that bulk mining by a modified block caving

system is possible and feasible. Special emphasis must be put on the

selection of large enough orebodies to justify a block caVing method.

Draw control must be carefully supervised and all personnel involved

in the opera tion must be trained to realize the consequences which

will result if the prescribed procedures are not followed.

Modifications may be made in operational procedure as the

mining engineer sees fit, but any changes must not interfere with the

sought-for end result, which is, to mine the ore on a scale large

enough and at a low enough unit cost per ton to keep mines on the

Gogebic Range in a favorable competetive position with producers from

other areas.

32

REFERENCES

(1) Report of Investigations No. 4155 December, 1947 - Investigation

of the Iron-Bearing Forll\9.tion of the Western Gogebic Range, Iron

County, Wisconsin - United States Department of the Interior ­

Bureau of Mines by Paul Zinner and Clyde L. Holmberg.

(2) The Geology of the Gogebic Iron Range of Wisconsin by H. R. Aldrich

Wisconsin Geolo.gical and Natural History Survey - Bulletin 71 ­

Economic Series No. 24 - 1929

(3) The Geology of the Lake Superior Region by Charles R. VanHise and

Charles K. Leith - United States Geological Survey - 1911.

(4) Mining Engineers' Handbook by Robert Peele - Vol. I

(5) Panel Caving at the Creighton Mine of the International Nickel

Company of Canada, Ltd. by A. E. Brock, R. J. McCormick and

W. J. Taylor Transactions of the Institute of Mining and Metal­

lurgy - Vol. 65, Part 2, 1955-56.

(6) Block Caving at the Kelley Mine, The Anaconda Company, Butte,

Montana, by C. C. Popoff - Bureau of Mines - Information Circular

No. 7758

(7) Block Caving Methods at the Sunrise Mine Platte County, Wyoming

by F. L. Wiedman - Bureau of Mines Information Circular 7759.

(8) Conversation with John Sharrer, District Engineer, Pickands

Mather Company, Ironwood, I1dchigan.

33

ACKNOWLEDGMENT

An expression of indebtedness is due Mr. John L. Sharrer,

District Engineer, Pickands Mather Ore Mining Company, Ironwood,

Michigan, and to his company for the opportunity of visiting under­

ground mines on the Gogebic Iron Range and for his most able guidance

during the course of this investigation.

I wish to express row appreciation to Professor L. D. Clark

of the Department of Mining and Metallurgy of the University of

Wisconsin for his helpful advice and assistance.

r ;7 ..

Date./

Title•

Name

35

APPROVAL

The foregoing t hes i s is hereby appr oved as a creditable

necessarily e ndorse or appr ove any statement made , opinion expr essed,

to be understood that by this appr oval t he undersigned does not

prerequisite to the degr e e fo r which i t has been submitted . I t is

purpose f or which i t is submitted.

study of an engineering sub ject , carried out and presented in a

manner s Ufficiently satisfactory to warrant its acceptance as a

or c onc l usion drawn the r ein, but appr ove s t he thesis only f or the